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John  S.Prcll 










Cidl  &  Mechanical  Engineer. 


New  York  and  London. 

Copyright,  1895, 


Copyright,  1908, 

BY  , 





The  Author 

Takes  great  pleasure  in  renewing  the 

Dedication  of  this  Book 

To  his  Friend 


of  New  York, 

President  of  the 

Copper  Queeist  Mining  Company. 





The  collection  of  papers  which  forms  this  book  was  mostly 
prepared  in  moments  stolen  from  more  active  professional  duties, 
and  must  consequently  lack  the  uniformity  and  completeness 
which  is  compatible  only  with  ample  leisure  and  freedom  from 
other  more  pressing  cares. 

It  has  been  my  intention  to  confine  myself  principally  to 
facts  gleaned  from  my  own  experience,  and  only  to  touch  upon 
theoretical  questions  when  essential  for  the  understanding  of 
practical  facts. 

As  the  items  of  cost,  both  of  construction  and  subsequent 
operation,  are  amongst  the  most  important  of  all  the  practical 
questions  that  face  the  originators  of  new  smelting  enterprises, 
and  as  these  are  virtually  unattainable  to  the  general  public,  I 
have  gone  into  these  figures  in  considerable  detail,  not  calculating 
expenses  as  they  appear  on  paper,  and  when  everything  is  run- 
ning smoothly,  but  giving  the  actual  results  of  building  on  a 
large  scale,  and  of  smelting  many  thousand  tons  of  ores  under 
varying  circumstances,  and  in  all  of  the  ordinary  kinds  of  fur- 

Owing  to  the  magnitude  of  the  subject,  I  found  it  impossible  to 
touch  upon  the  so-called  "Wet  Methods"  without  increasing  the 
size  and  consequent  cost  of  this  volume  to  an  extent  that  might 
probably  peril  its  circulation. 

The  author  desires  to  acknowledge  the  valuable  assistance  of 
Mr.  J.  E.  Mills,  in  connection  with  the  geology  of  the  Butte  min- 
ing district,  and  to  credit  Mr.  H.  M.  Howe  and  Mr.  A.  F.  Wendt 
with  the  use  he  has  made  of  their  papers  on  ''Copper  Smelting" 
and  on  "The  Pyrites  Deposits  of  the  Alleghanies." 

But,  above  all,  he  has  to  thank  Mr.  James  Douglas  for  a 
thorough  and  minute  revision  and  criticism  of  his  manuscript  just 
before  publication. 

E.  D.  P.,  Jr. 
WAi.ror.E,  Mass.,  June,   1887. 



Since  the  last  thorough  revision  of  this  work,  the  metallurgy  of 
copper  has  been  greatly  modified  by  the  success  and  general  intro- 
duction of  Automatic  Calcining  Furnaces;  by  the  rapid  and 
extraordinary  development  of  the  Copper  Bessemer  process;  by 
the  important  and  far-reaching  improvements  in  Blast  Furnaces 
and  Reverberatories;  and  perhaps,  above  all,  by  the  gradual 
dawning  of  the  idea  that,  because  Copper  is  worth  fifteen  times  as 
much  as  Iron,  it  is  not  absolutely  necessary  to  expend  fifteen  times 
as  much  money  in  handling  and  treating  its  ores. 

The  fusion  of  sulphide  ores  by  the  heat  generated  from  their 
own  oxidation  (Pyritic  Smelting)  has  also,  lately,  become  an 
accomplished  fact  in  several  diiferent  localities — though  how  im- 
portant it  is  to  be  as  a  process  pure  and  simple  can  scarcely  yet  be 

The  electrolytic  refining  of  pig  copper  has  become  such  an 
important  feature  in  many  American  and  European  works,  that 
it  has  been  thought  proper  to  include  a  chapter  on  this  method. 
I  have  been  most  fortunate  in  securing  the  assistance  of  Mr. 
Maurice  Barnett  of  Philadelphia,  whose  standing  and  long  prac- 
tical experience  well  qualify  him  to  write  with  authority  upon 
this  subject. 

All  these  radical  changes  have  made  it  necessary  to  re-write 
this  work,  and,  also,  to  add  very  considerably  to  its  size. 
Taking  advantage  of  the  break  caused  by  a  professional  trip 
to  Australasia,  I  have  spent  months  in  studying  European 
copper  practice,  in  order  to  glean  what  might  be  useful  in 
our  own  work.  The  preparation  of  this  edition,  with  its  numerous 
elaborate  working  drawings  and  plates,  has  occupied  considerably 
more  than  a  year,  and  even   then  would   have  been  impossible 


without  the  liearty  co-operation  of  most  of  the  copper  smelters  of 
the  United  States,  and  the  direct  assistance  of  several  metallur- 
gists who  have  written  for  me  on  special  subjects. 

Thus,  Mr.  Robert  Sticht,  of  Montana,  has  contributed  an  ex- 
haustive and  valuable  paper  on  Pyritic  Smelting.  Mr.  A.  H. 
Low,  of  Denver,  has  furnished  much  valuable  original  material  in 
relation  to  the  assaying  of  copper.  Mr.  H.  A.  Keller,  Superin- 
tendent of  the  Parrot  Silver  &  Copper  Company  of  Butte,  has 
collaborated  with  me  in  writing  the  chapter  on  bessenierizing, 
and  has  furnished  me  with  material  that  can  only  be  supplied  by 
a  specialist  in  that  department,  and  that  will  be  appreciated  by 
every  copper  smelter  in  this  country  and  in  Europe.  Mr.  I.  H. 
Cluttou,  of  the  Messrs.  Elliott's  Metal  Company,  Lim.,  South 
Wales,  has  kindly  furnished  me  with  a  detailed  description  of  the 
Iodide  Copper  assay,  as  practised  in  Great  Britain. 

It  is  impossible  for  me  to  even  enumerate  the  names  of  gentle- 
men in  Europe,  Australasia,  and  America,  who  have  given  me 
assistance  in  preparing  this  edition,  and  in  gathering  the  material 
for  the  same.  I  must,  howeverj  particularly  thank  Mr.  Christo- 
pher James,  of  Swansea,  Mr.  Richard  Pearce,  of  Argo,  Colorado, 
and  Mr.  C.  M.  Allen,  of  Butte,  Montana. 

I  also  gratefully  acknowledge  courtesy  and  assistance  from 
Messrs.  N.  P.  Hill,  W.  L.  Austin,  H.  A.  Vezin  and  M.  I.  lies,  of 
Denver,  Colorado,  and  wish  to  thank  the  Directors  of  the  Boston 
&  Colorado  Smelting  works,  and  the  Globe  Smelter.  Also  Messrs. 
E.  P.  Matthewson,  A.  S.  Dwight,  Carl  Eilers,  and  A.  Raht,  of 
Pueblo,  Colorado,  and  the  Directors  of  The  Pueblo  Smelting 
Company,  The  Colorado  Smelting  Company,  and  The  Philadel- 
phia Smelting  and  Refining  Company.  Also  Mr.  Franklin 
Ballon,  of  Leadville,  Colorado,  and  the  Directors  of  the  LaPlata 
Smelter  and  the  Arkansas  Valley  Smelter.  Also,  Mr.  Otto  Stalman, 
of  Salt  Lake,  Utah;  and  Messrs.W.  A.  Clark,  F.  A.  Heinze,  R.  G. 
Brown,  A.  II.  Wethey,  H.  Williams.  H.  C.  Bellinger,  Captain 
Palmer,  C.  W.  Parsons  and  0.  Szontagh,  of  Butte,  Montana,  and 
the  Directors  of  The  Colorado  SmeltiTig  &  Mining  Company,  The 
Parrot  Silver  &  Copper  Company,  The  Butte  &  Boston  Mining 
&  Smelting  Company,  and  The  Montana  Ore  Purchasing  Com- 
pany. Messrs.  Frank  Klepetko,  G.  M.  Hyams,  and  the  Directors 
of  The  Boston  &  Montana  Consolidated  Mining  &  Smelting  Com- 
pany, of  Great  Falls,  Montana.  Messrs.  H.  Thofehrn,  V. 
Ray,  and   the   Directors  of  The  Anaconda  Mining  Company  of 


Anaconda,  Montana.  Mr.  A.  R.  Meyer,  President  of  The  Kansas 
City  Consolidated  Smelting  &  Refining  Company,  of  Kansas  City, 
Missouri.  The  Mattliiessen  &  Hegeler  Zinc  Company,  of  LaSalle, 
Illinois.  Mr.  H.  F.  Brown,  of  Chicago.  Mr.  Titus  Ulke,  of  Wash- 
ington. Mr.  James  Douglas,  of  New  York,  President  of  the 
Copper  Queen  Mining  Company.  Also  the  late  Lord  Swansea, 
and  Messrs,  T.  D.  Nicholls  and  William  Terrill,  of  Swansea,  and 
Mr.  Gerard  B.  Blkiugtou,  of  Pembrey,  Wales.  My  best  thanks 
are  due  to  the  Directors  of  the  Rio  Tinto  and  the  Tharsis  Com- 
panies for  their  unbounded  hospitality  at  their  Spanish  mines,  and 
to  their  local  superintendents  and  other  officials.  Also  to  M. 
Carnot,  of  the  Ecole  des  Mines,  and  to  MM.  Martin  and  Fenelais, 
of  Paris.  Also  to  the  late  Professor  Stelzner,  to  Professors 
Richter  and  Weisbach,  and  to  the  Royal  Bureau  of  Mines  at  Frei- 
berg. Also  to  Herren  Bergnieister  Schroeder,  Hiittenmeister 
Steinbeck,  and  the  Directors  of  the  Mansfelder  Gewerkschaft. 
To  the  Humboldt  Machine  Company, o  f  Kalk-on-the-Rhine.  Also 
to  Messrs.  Bowes  Kelly,  Win.  Knox,  H.  H.  Sticbt,  and  the  Di- 
rectors of  the  Broken  Hill  Mining  Company,  of  New  South  Wales. 
To  the  Hon.  John  Henry  and  Mr.  G.  F.  Beardsley,  of  Tasmania; 
and  many  others. 

It  is  only  fitting  that  I  should  acknowledge  my  especial  indebt- 
edness to  Messrs.  Fraser  &  Chalmers,  of  Chicago,  whose  knowledge 
of,  and  intimate  relations  with,  almost  every  important  mining 
district  in  the  world,  has  enabled  them  to  aflEord  me  assistance  and 
information  that  has  been  of  the  greatest  value. 

E.  D.  P.,  Jr. 

Dorchester,  Mass.,  August,  1895. 



CHAPTER  I.— Copper  and  its  Ores 1-13 

Properties  of  Copper,  1.  Effect  of  Impurities,  2.  Tempered  Cop- 
per, 3.  Compounds  of  Copper  and  their  Reactions,  6.  Ores  of 
Copper,  Native  Copper,  Cuprite,  7.  Melaconite,  8.  Malachite, 
Azurite,  9.  Chalcopyrite,  10.  Chalcocite,  11.  Bornite,  Tetrahe- 
drite,  13. 

CHAPTER  II. — Distribution  of  the  Ores  of  Copper 14-27 

The  Atlantic  Coast  Beds,  14.  The  Lake  Superior  Deposits,  16. 
The  Deposits  of  the  Rocky  Mountains,  and  Sierra  Nevadas,  17. 
The  Butte  Mines,  18.  The  Arizona  Copper  Mines,  20.  The 
Clifton  District,  21.  The  Bisbee  District,  24.  The  Globe  District, 
25.     The  Black  Range  Copper  District,  27. 

CHAPTER  III.— The  Sampling  and  Assaying  op  Copper 28-74 

Sampling  Ores,  28.  Automatic  Samplers,  29.  Losses  in  Ship- 
ment, 38.  English  Weights  and  Samples,  39.  The  Assaying  of 
Copper,  41.  The  Electrolytic  Assay,  43.  The  Cyanide  Assay,  52. 
Low's  Modified  Cyanide  Assay,  56.  The  Iodide  Assay,  59.  The 
Colorimetric  Assay,  65.  The  Lake  Superior  Fire  Assay,  65.  The 
Determination  of  Gold  and  Silver  in  Copper  Furnace-material,  67. 
A  Method  for  Determining  Sulphur  in  Roasted  Ores,  71. 

CHAPTER  IV.— The  Chemistry  op  the  Calcining  Process 75-86 

Varieties  of  Roasting,  76.  Behavior  of  Sulphide  Ores  during 
Roasting,  76.  Degree  of  Roasting,  81.  Matte  Assay,  82.  Calcu- 
lation of  Roasted  Ore  for  Smelting  Mixture,  82.  Chemical  Reac- 
tions, 84.  Capacity  of  Calcining  Furnaces,  85.  Loss  of  Copper 
during  Roasting,  86. 

CHAPTER  V. — The  Preparation  of  Ores  for  Roasting 87-103- 

Classification  of  Roasting  Appliances,  87.  Best  Size  to  Break  Ores, 
88.  Production  of  Fines,  89.  Cost  of  Breaking  Ores  by  Machinery, 
92.  The  Breaking  of  Ore  by  Hand,  93.  Cost  of  Breaking  Ore  by 
Hand,    95.      Granulation    of    Mattes    by    Water,    97.      Crushing 



Machinery,  98.    Machines  for  Preparatory  Crushing,  Jaw-crushers, 
99.     Machines    for    Fine     Crushing;    Stamps,     Ball    Pulverizers, 
Chilian   Mills,  Multiple-jaw   Crushers,  100.     Cornish  Rolls,  101. 
Elevators,  103. 

CHAPTER  VI.— The  Roasting  op  Ores  in  Lump  Form 104-170 

Heap-roa.sting,  104.  Injurious  Effects  of  Heap-roasting,  105. 
Remedies,  106.  Selection  of  Site  for  Roast-yard,  107.  Preparation 
of  Roast-yard,  108.  Elevated  Track  for  Roast-yard,  110.  Size  of 
Roast-heaps,  113.  Construction  of  Roast-heaps,  115.  Proper  Use 
of  Fuel  in  Heap  roasting,  117.  Firing  the  Roast-heap,  119.  Man- 
agement of  Roast-heaps,  120.  Loss  of  Copper  by  Leaching,  123. 
Removing  of  the  Roasted  Heap,  126.  Heap  Matte,  128.  Costs  of 
Heap-roasting,  132.  Results  of  Heap-roasting,  133.  V-inethod  of 
Heap-roasting,  136.  The  Heap-roasting  of  Matte,  137.  Costs  of 
Heap-roasting  Matte,  140.  Stall  roasting,  140.  Open  Stalls,  141. 
Manufacture  of  Slag-brick,  142.  Arrangement  of  Stalls,  150. 
Management  of  Roast  .stalls,  153.  Results  of  Stall-roasting,  156. 
Cost  of  Erecting  Roast-stalls,  158.  The  Stall-roasting  of  Matte, 
163.     The  Roasting  of  Lump  Ores  in  Kilns,  166. 

CHAPTER  VII. — The  Roasting  of   Oiies  in  Pulverized  Condi- 
tion  171-199 

Classification  of  Roasting  Furnaces,  Shaft-furnaces,  171.  The 
Gerstenhofer  Furnace,  the  Hasenclever  Furnace,  the  Maletra  Fur- 
nace, 172.  The  Stetefeldt  Furnace,  Pelatan's  Stall,  173.  Hand 
Reverberatory  Calciners,  with  Open  Hearth.  174.  Construction  of 
Calciners,  177.  Calciner  Stacks,  186.  Cost  of  Calcining  Furnaces, 
192.  Cost  of  Calcining  in  Hand-reverberatories,  193.  Muffle  Cal- 
ciners, Revolving  Cylinders,  194.  Cylinders  with  Continuous 
Discharge,  195.     Cylinders  with  Intermittent  Discharge,  196. 

CHAPTER  VIII.— Atjtom.vtic  Reverberatory  Calciners. 200-223 

Classification  of  Automatic  Calciners,  the  O'Harra  Furnace,  200. 
The  AUen-O'Harras  at  Butte,  201.  The  Pearce  Turret  Furnace, 
205.  The  Improved  Spence  Calciner,  214.  The  Brown  Horseshoe 
Calciner,  218.  The  Spence  Automatic  Desulphurizer,  220.  The 
Matthiessen  &  Hegeler  Company's  Calciner,  222.  Calciners  with 
Movable  Hearth.  Blake's  Calciner,  223. 

CHAPTER  IX.— The  Smelting  of  Copper 224-235 

The  Object  of  Smelting,  Advantage  of  High  Rate  of  Concentration, 
224.  Products  of  Smelting,  Blister  Copper,  226.  Copper  Bottoms, 
227.  Matte,  228.  Speiss,  281.  Slags,  232.  Flue-dusi,  234. 
Classification  of  Smelting  Methods,  235. 

CONTENTS.  xiii 


CHAPTER  X. — The  Chemistry  op  the  Blastfurnace 236-249 

Distinctive  Ffiatures  of  tlie  Blast-furnace,  236.  Reactions  in  the 
Blast-furnace,  237.  Example  of  Calculating  a  Blast-furnace 
Charge,  240. 

CHAPTER  XI. — Blast-purnace  Smelting   (with    Carbonaceous 

Fuel) 250-319 

Modern    American   Copper    Blast-furnace,    251.       Advantages    of 
Water-jacket  Furnaces,   253.     Heat  Abstracted   by  Jacket- water, 
258.    Water-jacket  Blast-furnaces,  Cast-iron  Jackets,  260.  Wrought 
iron  Jackets,  263.     Herreshoff  Furnace,  266.     Verde  Water-jacket 
Furnace,  270.    Furnace  Bottoms,  271.     Blowing-in  Water-jackets, 
275.       Forehearths.   281.      Mansfeld   Process,   282.      Ureat   Falls 
Foreheartii,  285.     Herreshoff  Forehearth,  287.     Orford  Siphon-tap, 
294.     Mathevvson's  Matte-trap,  297.     Reverberatory    Forehearths, 
298.     Size   and   Shape    of     Blast-furnaces,     303.       Blast-furnaces 
Simply  for  Resmelting,  303.     Suggestion  for  Butte  Practice,  306. 
Blast-furnaces  for  Partial  Oxidation,  307.     The    Charging  of    the 
Blast-furnace,  309.     The  Handling  of  Blast  furnace  Products,  310. 
Keller   on  Slag-pots,''313.      Mechanical   Pan-conveyers  for   Slag, 
Granulation  of  Slag  by  Water,  317. 

CHAPTER  XII. — Blast-purn.vces  Constructed  op  Brick 320-342 

The  Orford  Raschette  Furnace,  320.  Smelting-in  Bottom.  326. 
Irregularities  in  Running,  327.  Intermittent  Running  of  Blast- 
furnaces, 333.  Repairs  of  Furnaces,  .338.  Estimate  of  Cost  of 
Large  Brick  Blast-furnace,  341. 

CHAPTER  XIII.— General  Remarks  on  Blast-purnace  Smelting. 343-371 

Capacity  of  Blast-furnaces,  343.  Use  of  Wood  in  Blast-furnaces, 
845.  Use  of  Bituminous  Coal  in  Blast-furnaces,  346.  Size  of  Ore- 
charges,  350.  Treatment  of  Fine  Ore  in  Blast-furnaces,  353. 
Effect  of  Fine  Ore  in  Diminishing  Capacity  of  Blast-furnaces,  358. 
Blowers  and  Accessory  Blast  Apparatus,  363.  Results  at  Butte 
Smelters,  366.  Cost  of  Smelting  in  Water- jacketed  Blast- 
furnaces, 368. 

CHAPTER  XIV.— Pyrittc   Smelting 372-395- 

Definition  of  Pyritic  Smelting,  372.  Cost  of  Coke  per  Ton  Ore 
Smelted  in  Ordinary  Smelting,  374.  Advantages  of  Pyritic  Smelt- 
ing. 376.  Effect  of  Shape  of  Furnace  on  Pyritic  Smelting,  379. 
Pyritic  Smelting  with  Column  Charging,  382.  Pyritic  Smelting 
with  Layer  Charging,  386.  Comparison  of  Chemical  Reactions  in 
Ordinary  Smelting  and  in  Pyritic  Smelting,  390.  Summary  for 
Pyritic  Smelting,  393. 



CHAPTER   XV. — Pyritic    Smelting — Its    History,   Principles, 

Scope,  Apparatus,  and  Practical  Results 396-441 

Distinction  between  true  "  Pyritic  Smelting  "  and  mere  Concentra- 
tion of  the  Gold  and  Silver  by  the  Use  of  Pyritic  Ores,  396.  His- 
tory of  Pyritic  Smelting,  400.  Hollway's  Experiments,  401. 
Pyritic  Smelting  at  Toston,  Montana,  411.  Austin's  Patents,  412. 
Boulder  Pyritic  Smelter,  Bi  metallic  Pyritic  Smelter,  415.  Table 
of  Pyritic  Furnaces  in  the  United  States,  Literature  of  the  Process. 
416.  Apparatus,  and  Principal  Scope  of  the  Process,  417.  Recov- 
eries of  Values  in  Pyritic  Smelting,  427.  The  Cost  of  Pyritic 
Smelting,  435.     Cost  of  Works  for  Pyritic  Smelting,  438. 

CHAPTER  XVI.— Reverberator Y   Fur>-aces 44^-527 

The  Chemistry  of  Reverberatory  Smelting,  442.  The  Evolution  of 
the  Modern  Reverberatory,  445.  Reverberatory  Practice  at  Butte, 
Montana,  451.  Anaconda  Hot-air  Reverberatory,  454.  The  Butte 
Reverberatories,  457.  Modern  Reverberatories,  458.  Brick 
Bottoms,  466.  Hot  Air  in  Reverberatory  Practice,  Depth  of  Bottom 
below  Skimming  Doors,  467.  Handling  Slag  and  Matte,  471. 
Labor  on  Reverberatories,  474.  Dust-chambers,  475.  Construc- 
tion of  Reverberatory  Smelting-furnaces.  476.  Chimneys,  481. 
Reverberatory  Hearths,  484.  Management  of  Furnace,  487.  Cost 
of  Running  a  Reverberatory  Furnace,  490.  Cost  of  Erecting  a 
Modern  Reverberatory  Furoace,  492.  Smelting  for  White  Metal, 
493.  The  Making  of  Blister  Copper,  495.  Copper  Refining,  499. 
Cost  of  Copper  Refining,  509.  Refining  Copper  with  Gas,  509. 
The  Blister  Process  at  Atvidaberg,  510.  The  Refining  Process  at 
Atvidaberg,  Gas  Furnaces  in  America,  518.  The  "Direct  Method  " 
of  Copper  Refining,  519. 

CHAPTER  XVII.— The  Bessemerizing  of  Copper  Mattes 528-575 

Table  of  American  Converters,  529.  Peculiarities  of  Copper  Bes- 
semerizing,  534.  Materials  Suitable  for  Bessemerizing,  538, 
Description  of  a  Converter  Plant,  .540.  Cost  of  a  Three-converter 
Plant,  548.  Converter  Practice,  549.  Results  of  Bessemerizing 
Mattes.  551.  Remelting  Cupola,  553.  Converter  Reactions,  557. 
Indications  Furnished  by  Converter  Flame,  558.  Time  Required  for 
Converting  Matte,  563.  Labor  Required  at  Converters,  564.  Silver 
and  Gold  in  Copper  Bars,  565.  Metallurgical  Losses  in  Bessemer- 
izing, 566.  Cost  of  Converting  Copper  Matte,  569.  Converter 
Linings,  570. 

CHAPTER  XVIII.— The  Electrolytic  Refixixg  of  Copper 576-606 

Refining  by  "  Multiple  Arc."  and  by  "  Series  "  Methods.  577.  Com- 
parison of  Costs  in  these  Two  Methods,  581.    Boilers  and  Engines  for 



Electrolytic  Work,  583.  Generators,  584.  Anodes,  Cathodes,  and 
Tanks,  586.  Conductors,  Electrolyte,  etc.,  594.  Current 
Required,  596.  Treatment  of  tlie  Gold  and  Silver-slimes,  602. 
Cost  of  Treating  Slimes,  604.     Cost  of  Electrolytic  Refinery,  606. 

CHAPTER  XIX. — Selection   of   Process   and  Arrangement  of 

Plant 607-«28 

The  Three  Groups  of  Ores  to  be  Treated.  Native  Copper  Mines,  607. 
Oxidized  Ores,  608.  Sulphide  Ores,  610.  Choice  of  Location  and 
Site  of  Smelter,  618.  Railway  Tracks.  620.  Handling  Slag  and 
Matte,  621 .  Foundations  and  General  Construction  of  Works,  623. 
Transportation  of  Material  in  the  Works,  Water  Supply,  624. 
Supplies,  627.     Management  of  Workmen,  628. 

Unless  otherwise  specified,  the  "  short  ton  "  of  2,000  pounds  (907.2  kilos)  is 
used  in  this  work. 

One  pound  (avoirdupois) ....  0.4536  kilos. 

One  foot  =  12  inches 0.3048  metres. 

One  gallon 3.7854  litres. 

One  ounce  (Troy  weight,  as  used  for  precious  metals)  31.1  grams. 

To  reduce  "  ounces  per  ton  of  2,000  pounds"  to  per  cent.,  multiply  by 
0.00343.  Example  :  What  is  the  value,  expressed  in  per  cent.,  of  an  ore  con- 
taining 155  ounces  silver  per  ton  ? 

Answer— 155  X  0.00343  =  0.53165  per  cent. 

To  change  per  cent,  into  ounces  per  ton  of  2,000  pounds,  multiply  by  292. 
Example  :  An  ore  contains  0.01543  per  cent,  of  gold,  how  much  is  this  per  ton 
of  2,000  pounds  when  expressed  in  ounces  ? 

Answer — 0.01543  X  292  =-  4.5  ounces  per  ton. 

The  common  measure  of  wood  (fuel)  in  the  United  States  is  the  "  cord  "  of 
128  cubic  feet. 

An  American   dollar  contains  one   hundred  cents,  and  is  equal  to  about 
4  shillings    2  pence English,  or 

4  marks       16  pfennige German,  or 

5  francs      21  centimes French. 

The  value  of  certain  foreign  coins  in  American  money  has  been  determined 
by  the  United  States  Treasury  Department  as  follows: 


Title  of  Coin. 

Value  in  U.  S. 



f 0.72.7 
0  54  6 



Chili    ...      . 


0  91  2 

Cuba    .... 


0  92  6 



0  19  3 

Germany , 



(Treat  Britain 







Iji  ra   

0  19.3 


Yen  (gold) 


Yen  (silver) 

0  78.4 













Portugal. . . 







Sweden   .  . 



0  26.8 




Copper  is  a  red  metal,  having,  when  pure  and  uou-porons,  a 
specific  gravity  of  8.945  in  vacuo  and  at  the  freezing  point  of 

The  slight  porosity  of  ordinary  commercial  copper  reduces  this 
figure  to  8.15 — 8.6. 

The  capacity  of  copper  for  conducting  heat  stands  very  high, 
being  898  when  gold  is  called  1000. 

Its  electrical  conductivity  is  931,  silver  being  1000. 

"When  copper  is  heated  to  within  200  or  300  degrees  of  its  melt- 
ing point,  it  becomes  brittle  and  friable,  and  may  be  actually 

The  melting  points  of  substances  that  are  not  easily  fusible  have 
not  been  determined  with  exactness.  As  an  approximation,  copper 
may  be  assumed  to  melt  at  3000  degrees  Fahr.  (1093  degrees 

When  fused  it  possesses  a  sea-green  color.  It  is  volatile  when 
exposed  to  our  highest  attainable  temperatures,  such  as  the  electric 
arc,  or  the  oxy-hydrogen  flame. 

Molten  copper  has  the  property  of  absorbing  certain  gases,  such 
as  hydrogen,  carbonic  oxide,  sulphurous  acid,  etc.,  which  it  sets 
free  again  on  solidifying.  This  causes  serious  difficulties  in  mak- 
ing copper  castings,  and  requires  the  employment  of  special  pre- 
cautions to  prevent  porosity. 

The  malleability,   ductility,   softness,    and    strength    of    copper 

*  W.  Hainpe,  Zeitschrift  fur  Berg-  Hutten-  unci  SalineMtoesen,  1873,  p.  218 


are  extraordinarily  atfected  by  the  presence  of  minute  quanti'J(i:i 
of  certain  other  substances. 

Cuprous  oxide  dissolves  rapidly  and  homogeneously  in  metallic 
copper,  and,  according  to  Hampe's  researches,  produces  absolutely 
no  effect  until  present  to  the  extent  of  0.5  per  cent.  Even  1  ])vr 
cent,  of  this  substance  produces  but  a  very  slight  diminution  in 
toughness.  In  greater  quantities,  it  lowers  the  ductility  of  the 
metal,  but  may  be  present  up  to  8  per  cent,  to  18  per  cent,  without 
rendering  the  copper  untit  for  most  purposes. 

Iron,  when  present,  is  generally  distributed  through  the  copper 
with  much  regularity.  Reliable  data  as  to  the  effect  of  this  sub- 
stance on  the  tensile  strength  and  electrical  conductivity  of  copper 
are  not  to  be  had.  But  the  weight  of  evidence  seems  to  show  that 
the  minute  proportion  of  iron  so  frequently  present  in  refined  cop- 
per has  no  appreciable  effect  on  any  of  its  useful  qualities,  except, 
possibly,  its  electrical  conductivity. 

Z/;/c  forms  an  alloy  with  copper  in  every  proportion,  and  slightly 
lowers  its  ductility  at  high  temperatures.  But  until  the  zinc 
reaches  at  least  18  per  cent.,  the  tenacity  of  the  alloy  at  ordinary 
temperatures  does  not  seem  to  be  lessened. 

Lead,  while  a  useful  addition  to  copper  that  is  to  be  employed 
for  certain  meclianical  purposes,  and  assisting  in  procuring  solid 
castings,  has  a  decidedly  injurious  effect  if  present  in  too  large 
quantities.  One-third  of  one  per  cent,  is  sufficient  to  make  the 
metal  both  red-short  aud  cold-short,  while  0.75  per  cent,  will  ruin 
copper  for  any  ordinary  purpose. 

Tin  also  alloys  with  copper  in  all  proportions,  and  begins  to 
injure  its  ductility  wlien  present  in  quantities  of  1  per  cent. 

yickel  is  frequently  present  in  commercial  copper,  and  up  to 
0.3  per  cent,  seems  to  produce  no  injurious  results. 

Bismuilt  is,  perhaps,  the  worst  enemy  of  the  copper  refiner; 
for,  in  spite  of  its  oxidizability,  it  clings  to  copper  with  much 
tenacity,  and  affects  its  properties  in  the  most  surprising  manner. 
Hampe  finds  that  merely  0.02  per  cent,  is  sufficient  to  make  the 
metal  distinctly  red-short,  and  cold-shortness  begins  with  0.05  per 
cent,  bismuth. 

Arsenic  is  not  so  injurious  to  copper  as  is  often  supposed. 
Hampe  cannot  find  that  0.5  per  cent,  of  this  substance  has  any 
bad  effect  on  copper,  except  to  diminish  its  electrical  conductivity; 
and  copper  containing  0.8  per  cent,  arsenic  was  drawn  by  him  into 
the  finest  wire. 


Antimony  up  to  0.5  j)er  cent,  acts  very  much  like  arsenic,  and 
the  lovveriug  of  the  metal's  electrical  conductivity  does  not  seem 
to  be  accompanied  with  a  loss  of  either  strength  or  ductility.  So 
Hampe  decides  after  a  long  series  of  most  careful  researches. 
When  present  in  quantities  greater  than  0.5  j)er  cent.,  antimony  is 
much  more  injurious  than  arsenic. 

lellurium,  which  is  by  no  means  a  rare  constituent  of  copper  in 
certain  districts,  produces  red-shortness,  even  though  present  in 
very  small  amounts.  But  Mouchel  claims  that  at  ordinary  tem- 
peratures, the  addition  of  0.1  percent,  of  tellurium  largely  increases 
the  tensile  strength  of  copper  without  materially  lessening  its 

Silicon  has  been  very  thoroughly  studied  by  Hampe  as  to  its 
effect  on  copper.* 

The  addition  of  0.5  per  cent,  of  this  metal  causes  a  distinct  low- 
ering of  electrical  conductivity.  But  copper  may  contain  3  per 
cent,  of  silicon  before  its  toughness  or  malleability  is  affected.  Six 
per  cent,  makes  copper  brittle,  increasing  as  the  silicon  is  increased, 
until  at  11.7  per  cent,  the  alloy  is  as  brittle  as  glass. 

Sulphur  in  quantities  of  0.25  per  cent,  lowers  the  malleability 
of  copper.  One-half  of  per  ctjnt.  jirodnces  cold-shortness,  though 
curiously  enough  such  copper  is  not  red-short. 

Phosphorus  in  small  quantities  seems  to  produce  no  injurious 
effect.     At  0.4  per  cent,  red-shortness  is  developed. 

Carbon  is  not  at  all  taken  up  by  molten  copper,  according  to 
Hampe's  late  researches,  though  not  long  ago  it  was  believed  to  be 
absorbed  in  considerable  quantities,  and  to  affect  the  copper  most 
seriously.  Too  long-continued  poling  was  thought  to  bring  about 
this  result,  but  it  seems  to  be  now  conclusively  established  that 
the  evil  effect  of  "overpoling"  copper  in  the  refining  furnace  is 
due  to  the  reduction  of  certain  metallic  compounds,  which,  when 
present  in  the  copper  in  an  oxidized  condition,  produce  no  visible 
effect.  Such  compounds  are  arsenates  and  antimonates  of  lead 
and  bismuth,  mixtures  of  oxides  of  lead  and  copper,  etc.  The 
gases  arising  from  poling,  such  as  carbonic  oxide,  hydrogen,  etc., 
may  also  be  absorbed  by  the  copper,  and  affect  it  injuriously. 

Tempered  copper  has  been  put  upon  the  market  by  the  Eureka 
Tempered  Copper  Company,  samples  of  which  were  examined  at 
the  Versuchsanstalt  fiir  Bau-und  Maschinen  Material  with  the 
following  results,  the  investigation  having  been  made  by  P.  Kirsch : 

*Chemiker  Zcititng.  1892,  No.  42. 



Ordinary  Copper. 
Per  cent. 

Tempered  Copper. 
Per  cent. 




















As  ■will  be  seen  from  the  foregoing  analyses,  the  ditference  of 
tempered  copper  from  copper  of  ordinary  commercial  quality,  as 
far  as  its  composition  is  concerned,  is  but;  slight. 

The  coppers  of  which  the  analyses  are  given  above  were  mechan- 
ically tested,  with  the  following  resnlts: 


Strength  in       Elastic  Limit  Contraction  in 

1  kgs.*    per  sq.        in  kgs.  per       intension.  Area, 

mm.  sq.  mm.  "^''  cent.  p^^   cent. 

Tension,  tempered 

Tension,  Tempered 

Tension,  unteiupered | 

'i'ension.  nnteiiipered..    ..' 
Compression,  tempered... 
Compression,  tempered. .. 
Compression,  untempered 
Compression,   untempered  I 

19.. 58 




*1  kg.  per  square  mm.  =•  1.425.4.i  pounds  per  square  inch. 

The  tests  and  analyses  qnoted  above  were  carried  out  in  America, 
and  are  quoted  for  the  sake  of  comparison  with  those  performed 
at  the  Versnchsanstalt,  which  were  as  follows: 

(a)  Modulus  of  Elasticity. — The  modulus  of  elasticity  deter- 
mined on  a  specimen  tested  in  tension  was  10,000  kilos  per  square 
mm.  The  modulus  determined  by  compression  tests  was  2,930 
kilos  per  square  mm.,  with  a  load  of  2.-5  kilos  per  square  mm., 
and  1,0"20  kilos  per  square  mm.  with  a  load  of  T.2  kilos  per 
square  mm. 

{h)  Tensile  Strength.— Test  pieces  used:  Sheet,  0.11  mm.  in 
thickness,  oO.'i  kilos  per  square  mm.;  sheet,  0.13  mm.  in  thick- 

COPPEK    AND    ITS    OKES.  5 

ness,  G?.9;  shee^,  0.55  mm.  iu  thickness,  56.8;  sheet,  0.04  mm. 
in  thickness,  53.4;  sheet,  1.19  mm.  iu  thickness,  52.3;  wire,  0.50 
mm.  iu  diameter,  31.8;  wire,  0.80  mm.  in  diameter,  72.0;  wire, 
1.(55  mm.  in  diameter,  52.0;  wire,  2.60  mm.  in  diameter,  50.0; 
wire,  4.20  mm,  in  diameter,  47. G;  rod,  87  mm.  iu  diameter,  19.0. 

The  last  named  specimen  had  an  elastic  limit  of  8.1  kilos  per 
square  mm.  A  compression  test  was  made  iu  which  deformation 
began  when  the  load  had  rea(!hed  8.1  kilos  per  square  mm.  The 
load  could  be  increased  to  219  kilos  per  square  mm.  without  pro- 
ducing cracks,  although  the  test  piece,  which  was  originally  30 
mm.  in  height,  had  been  shortened  to  7.8  mm. 

(c)  Ductililty. — The  extension  given  by  the  sheet  varied  between 
0.2 — 2.0  per  cent.,  while  that  of  the  wire  was  0.1 — 0.2  per  cent, 
and  that  of  the  rod  13.1  per  cent.,  while  the  contraction  of  area  at 
the  point  of  fracture  of  the  latter  was  33  per  cent.  From  these 
tests,  as  well  as  by  winding  tests  with  the  wire,  it  appears  that  the 
material  possesses  great  ductility. 

The  foregoing  series  of  tests  shows  that  tempered  copper  possesses 
properties  that  distinguish  it  from  the  ordinary  material,  its 
strength  in  pieces  of  small  section  being  noticeably  high,  although 
that  of  larger  test  pieces  is  by  no  means  remarkable,  as  it  shows 
the  tensile  strength  of  only  10  kilos  per  square  mm.,  Avhile  ordi- 
nary commercial  copper  gives  20—25  kilos  per  square  mm.  Cast- 
ings made  of  it  are  of  good  quality,  and  its  electrical  conductivity 
is  high.* 

Copper  is  easily  soluble  iu  nitric  acid,  in  aqua  regia,  and  in 
strong  boiling  sulphuric  acid. 

In  dilute  sulphuric  and  muriatic  acids,  with  admission  of  air,  it 
dissolves  slowly. 

The  metallurgical  processes  for  obtaining  copper  from  the 
greater  proportion  of  its  ores  are  based  upon  its  strong  affinity  for 
sulphur,  wherein  it  exceeds  every  other  metal. 

If  an  impure  plate  of  copper  be  used  as  an  anode  in  an  acid 
solution  of  sulphate  of  copp  r,  and  another  plate  be  taken  as 
cathode,  a  properly  regulated  electric  current  will  precipitate 
chemically  pure  copper  upon  the  cathode,  the  anode  dissolving  in 
the  same  ratio,  while  the  impurities  contained  in  the  latter  (in- 
cluding often  gold  and  silver),  remain  as  a  residual  mud.     This  is 

*  Mittheil.  Teclm.  Oetcerhe- Museums,  1891,  261-267,  through  Journal  of  the 
Society  of  Chemical  Industry  for  February,  1892. 


tlic  basis  of  the  modern  electrolytic  refining  of  coi)per,  first  made 
a  technical  and  commercial  success  by  t!ie  Messrs.  Elkington  of 

Compounds  of   Copper   Most   Important  to   ^Ietallurgists, 

AND    their    EeACTIONS. 

Copper  has  two  oxides. 

Cuprous  oxide  melts  at  a  red  heat  without  decomposition.  It 
is  freely  dissolved  in  molten  metallic  copper.  In  a  pulverized 
condition  it  can  easily  be  changed  into  the  higher  oxide  by  gently 
heating  it  in  air.  It  is  easily  reducible  to  metallic  copper,  and 
forms  a  fusible  slag  with  silica. 

Melted  with  subsulphide  of  copper  in  proper  proportions,  the 
copper  in  both  substances  is  reduced  entirely  to  the  metallic  state, 
while  the  sulphur  of  the  matte  combines  witli  the  oxygen  of  the 
cuprous  oxide,  and  forms  sulphurous  acid  gas.  This  reaction, and 
a  few  analogous  ones  that  follow,  are  the  main  basis  of  our  treat- 
ment of  most  ores  of  copper. 

Subjected  to  the  same  treatment  with  sulphide  of  iron,  the 
copper  combines  with  sulphur  to  form  a  subsulphide,  which,  to- 
gether with  a  portion  of  the  undecomposed  sulphide  of  iron,  forms 
ft  matte,  while  the  iron  that  has  been  changed  to  a  protoxide  by 
the  oxygen  of  thecujirous  oxide,  combines  with  silica,  that  should 
also  be  present  to  form  slag.  The  matte  being  heavier  sinks  to 
the  bottom  and  can  easily  be  separated  from  the  supernatant  slag, 
and  the  first  step  in  the  fusion  of  copper  ores  has  been  accomplished. 

Cupvic  oxide  is  infusible.  It  is  easily  reduced  to  metal  by 
hydrogen  or  carbonic  oxide  gases. 

It  will  not  combine  with  silii^a  to  form  a  slag  at  ordinary  metal- 
lurgical temperatures  unless  there  be  such  conditions  present  that 
it  is  mostly  reduced  to  cuprous  oxide. 

Its  reactions  with  sulphide  of  iron  and  silica  lead  to  the  same 
result  as  in  the  case  of  cuprous  oxide.  That  is,  we  obtain  a  matte, 
containing  the  copper  as  a  subsulphide  together  with  indeterminate 
sulphides  of  iron,  and  a  silicate  of  the  ferrous  oxide  that  has  been 
formed,  as  already  explained. 

Both  of  these  oxides  are  soluble  in  ammonia. 

Silicate  of  copper,  in  the  presence  of  a  strong  base,  such  as  fer- 
rous oxide  or  lime,  is  reduced  by  carbon  to  the  metallic  state. 

When  heated  with  sulphide  of  iron  we  obtain  sulphide  of  copper 
and  silicate  of  iron. 

COPPER   AND    ITS    ORES.  7 

Wlieu  heated  with  metallic  iron  we  obtaiu  silicate  of  iron  and 
metallic  copper. 

Sidphides  of  copper. — We  know  two  sulphides  of  copper:  CuS 
and  CiisS.  Only  the  latter  is  of  importance  metallurgicaliy,  as  the 
CuS  loses  one-half  its  sulphur  at  a  comparatively  low  temperature. 

This  CugS,  or  subsulphide  of  copper  (cuprous  sulphide),  melts 
at  a  low  temperature,  and  fuses  together  with  sulphide  of  iron  to 
form  an  apparently  homogeneous  substance  called  matte.  This  is 
the  object  of  the  first  smelting  operation  with  ordinary  ores. 
Much  of  the  sulphide  of  iron  contained  in  the  ore,  as  well  as  tlie 
sulphides  of  copper,  has  been  changed  into  oxides  by  a  preceding 
calcination  at  a  low  temperature.  The  reactions  described  al)ove 
having  taken  place,  we  obtain  a  matte  containing  all  of  the  copper 
as  a  subsulphide  with  much  sulphide  of  iron,  while  the  alkaline 
earths  and  the  protoxide  of  iron  combine  to  a  worthless  slag. 

Under  favorable  circumstances,  and  with  low-grade  ores,  we 
may  thus  sometimes  diminish  our  material  to  be  treated  by  90 
per  cent,  or  more  in  this  single  fusion,  and  thus  obtain  practically 
all  of  the  copper  (as  well  as  any  gold  or  silver  present),  in  a  very 
small  quantity  of  matte,  on  which  we  can  afford  to  put  considera- 
ble expense. 

The   Ores   of   Copper. 

Although  the  copper-bearing  minerals  are  numerous,  yet  those 
of  commercial  importance  are  few  in  number,  and,  for  the  most 
part,  quite  simple  in  chemical  composition.  The  following  min- 
erals may  be  properly  considered  ores  of  copper,  and  are  found  in 
the  United  States  in  the  localities  enumerated. 

Native   Metallic   Copper. 

Aside  from  the  extensive  occurrence  of  this  metal  in  the  Lake 
Superior  region  and,  more  sparingly,  at  Santa  Rita,  New  Mexico, 
it  is  found  very  frequently  as  a  product  of  decomposition,  though 
seldom  in  sufficient  quantities  to  render  it  of  any  commercial 
importance.     It  is  usually  remarkable  for  its  purity. 

Cuprite,  or  Red  Oxide  of  Copper.  CujO;  88.8  Cu,  11.2  0. 

This  mineral  occurs  solely  as  a  product  of  decomposition,  and 
while  quite  widely  distributed,  is  nowhere  an  ore  of  any  impor- 
tance, except  in  the  Southwestern  carbonate  mines,  where  it  some- 


times  permeates  large  masses  of  iron  oxide,  notably  increasing  tbeir 
copper  contents.  Quite  large  lumps  of  this  mineral  are  found  in 
the  Santa  Rita  miues,  and  are  evidently  the  result  of  an  oxidation 
of  nodules  of  metallic  copper,  the  unaltered  center  being  usually 
preserved  of  greater  or  less  size.*  Many  of  the  Butte  City  veins, 
as  well  as  fissures  throughout  the  Eastern  Coast  Range,  carry  this 
mineral  in  their  upper  portions  as  a  product  of  the  decomposition 
of  sulphide  ores. 

Melacoxite,  Black  Oxide  of  Copper,  CcO;  79.8  Cr,  ^0.2  0. 

This  ore,  with  its  metallic  contents  usually  in  part  replaced  by 
oxides  of  iron  and  manganese,  is  not  quite  so  widely  distributed  as 
the  sub-oxide,  but  is  more  frequently  found  in  masses  sufficiently 
large  to  pay  for  extraction.  Its  most  remarkable  occurrence  in 
the  I.^nited  States  was  in  the  Blue  Ridge  mines  of  Tennessee, 
North  Carolina,  and  Virginia,  where  the  upper  portion  of  the 
beds  furnished  a  very  large  amount  of  from  2l)  to  50  per  cent, 
ore.  Laving  the  appearance  of  melaconite,  and  giving  rise  to  expec- 
tations that  were  always  shattered  after  passing  through  this  rich 
zone  and  reaching  the  lean,  unaltered  pyrites  below.  This  so- 
called  black  oxido  of  the  Blue  Ridge  regionf  seems  to  be  an  inti- 

*An  average  sample  of  thirteen  tons  of  concentrates,  taken  by  the  author  at 
Sanla  Rita,  in  1S81.  and  partially  analyzed  under  his  supervision,  gave,  after 
continuing  the  concentration  by  hand  to  almost  complete  removal  of  the  rock 
constituents  : 

Oxides  of  copper 13.42 

Carbonates  of   copper 1.27 

Oxides  of  iron 0. 13 

Metallic  iron  (from  stamps) 0.29 

Sulphur 0.11 

Insoluble  residue 0.37 

Metallic  copper 83.66 

Zn,  Ag.  Co.  Ni,  Pb.  Mn Traces 


This  analysis  presented  points  of  considerable  difficulty,  especially  in  deter 
mining  the  amount  of  oxide  of  copper  in  tlie  presence  of  metallic  copper. 
Entirely  satisfactory  results  were  not  obtained  :  but  the  method  proposed  by 
W.  Hampe.  by  means  of  nitrate  of  silver,  yielded  the  only  figures  that  could 
lay  the  slightest  claims  to  accuracy.  Since  this  was  written,  the  combustion 
method  has  been  so  perfected  that  it  will  undoubtedly  take  the  place  of 
Hampe's  process  in  determining  the  amount  of  oxygen  in  the  presence  of 
metallic  copper. 

^  PyriteA  Deposits  oftlie  Alleghanies,  by  A.  F.  Wendt. 

COPPER    AND    ITS   ORES.  9 

mate  mixture  of  glauce,  oxide,  carbonate,  and  sometimes  finely 
divided  native  copper.  Two  analyses,  by  Dr.  A.  Trippel,  show 
their  constituents: 

Oxide  of  copper 5.75  3.80 

Sesquioxide  of  iron 1.50  63 

Sulphur 18.75  25.40 

Copper 71.91  41.00 

Iron 93  26.56 

Soluble  sulphates  of  copper  and  iron 72  178 

99.56  99.17 

A  pile  of  such  ore,  laid  on  a  bed  of  cord  wood  and  moistened, 
often  ignites  the  wood  below,  and  thus  roasts  itself  without  firing. 

Malachite,  CuCOj  +  Cu  (OH)^;  71.9   CuO,  19.9  CO.,.  8.2  HO. 

This  is  a  much  more  valuable  compound  of  copper  than  the  two 
preceding  oxides,  from  a  commercial  standpoint;  although  no 
mines  in  the  United  States  furnish  malachite  of  sufficient  purity 
to  fit  it  for  ornamental  purposes. 

While  it  may  be  said  to  occur  in  widely  distributed  but  ordi- 
narily in  non-paying  quantities,  in  the  upper  decomposed  regions 
of  most  copper  deposits,  there  are  certain  localities  in  which  it 
forms  the  priuci])al  ore  of  this  metal.  It  is  very  seldom  found  in 
a  state  of  purity,  but  is  mixed  with  various  salts  of  lime  and  mag- 
nesia, oxides  of  iron  and  manganese,  silica  in  its  various  forms  of 
quartz,  chalcedony,  flint,  chert,  and  jasper,  and,  when  seemingly 
present  in  large  quantities,  it  often  forms  only  worthless  incrus- 
tations, or  merely  colors  nodules  and  masses  of  valueless 
material.  It  is  then  difficult,  and  in  some  cases  impossible,  to 
form  any  accurate  opinion  of  the  tenor  of  the  ore  from  its  external 
appearance.     It  is  an  important  ore  in  the  Southwest. 

AzuRiTE,  3  CuCOs  +  Cu  (0H)2;    69.2   CuO,  25.6    OOo,  5.2  HO. 

This  mineral  requires  only  a  passing  notice.  It  is  distributed 
in  the  same  manner  and  occurs  under  the  same  conditions  as  its 
sister  carbonate,  but  in  very  much  smaller  amounts.  It  occurs  in 
profitable  quantities  only  in  some  of  the  Southwestern  mines. 
Specimens  of  this  mineral  are  found  with  malachite  and  calc-spar 
in  the  Longfellow  mine,  exceeding  in  beauty  anything  of  the  kind 
that  is  known  elsewhere  in  the  United  States. 

lU  ilODEliX    COl'i'Eli   SMELTING. 

Chalcopyrite,  CUgS,  FE2S3 ;   34.4  Cu,  30.5   Fe,  35.1   S. 

This  is  by  far  the  most  widely  diatributed  ore  of  this  metal,  and 
furnishes  the  greater  proportion  of  the  world's  copper.  It  occurs 
principally  in  the  older  crystalline  rocks,  frequently  accompanied 
with  an  overwhelming  percentage  of  iron  pyrites,  in  bedded  veini^, 
in  Newfoundland,  in  Quebec,  Canada,  in  Vermont,  Virginia, 
Georgia,  Tennessee,  and  Alabama. 

The  value  of  copper-bearing  fissure  veins  below  the  lin)it  of  sur- 
face decomposition  is  nearly  always  due  to  this  mineral.  In  some 
localities  the  chalcopyrite  forms  with  pyrite  a  fine-grained  mechan- 
ical mixture,  varying  in  color  with  its  percentage  of  copper  from 
deep  yellovv  to  steel-gray.  This  substance  is  easily  recognized 
under  the  microscope  as  a  mechanical  mixture,  and  not  a  chemical 
compound.  In  most  of  the  carbonate  mines  of  the  Southwest  that 
have  attained  any  considerable  depth,  chalcopyrite  is  already  be- 
coming apparent,  in  minute  specks;  and  it  is  highly  probable 
that  the  altered  ores  near  the  surface,  with  their  valuable  admix- 
ture of  ferric  oxides,  are  all  due  to  the  decomposition  of  this  min- 
eral. The  sulphureted  fissure-veins  of  the  Kocky  Mountair.s  and 
Sierra  Nevada  are  seldom  free  from  this  mineral,  although  their 
value  almost  invariably  depends  upon  their  precious  metal  contents. 
The  remarkable  purple  ores  and  copper  glance  of  Butte  City, 
Montana,  have  already  in  several  mines  given  place  in  depth  to  the 
universal  yellow  sulphide. 

In  the  vast  bodies  of  bisulphide  of  iron  (iron  pyrites),  that  fur- 
nish so  large  a  proportion  of  the  material  for  the  world's  manufac- 
ture of  sulphuric  acid,  copper  pyrites  is  frequently  present  in. 
paying  quantities. 

Silver  and  gold  are  also  commonly  present  in  minute  amounts, 
and  one  of  the  most  interesting  feats  of  metallurgical  chemistry  is 
tlie  profitable  extraction  of  these  metals  from  ores  carrying  only 
3  per  cent,  copper,  less  than  one  ounce  of  silver,  and  four  or  five 
grains  of  gold.  (It  is  a  curious  fact  that  monosulphide  of  iron, 
though  often  rich  in  copper  pyrites,  and  noticeably  so  in  nickel, 
very  rarely  carries  more  than  the  merest  traces  of  the  precious 

In  these  great  pyrites  deposits,  a  concentration  of  the  silver 
frequently  takes  place  just  at  the  line  of  junction  between  the 
oxidized  surface  gossan  and  the  unaltered  pyrites  below. 

At  one  of  the  surface  openings  on  the  Rio  Tinto  deposit,  the 

COPPER    AND    ITS    ORES.  11 

line  of  demarcatiou  between  the  oxidized  and  sulphide  ores  is  very 
sharp.  The  gossan  at  this  place  was  about  sixty  feet  thick,  and. 
just  below  this  capping  of  red  iron  ore,  and  resting  upon  the 
almost  unaltered  pyrites,  was  a  soft,  grayish  earthy  deposit  from 
half  an  inch  to  six  inches  in  thickness,  and  containing  from  50  to 
150  ounces  silver  (0.17  per  cent,  to  0.52  per  cent.)  per  ton,  while 
the  original  pyrites  contained  probably  about  one  ounce,  or  less,  to 
the  ton.  A  sample  that  I  took  from  many  spots  assayed  6G  ounces 
(0.33  per  cent.)  silver  and  9  per  cent.  lead. 

I  had  the  pleasure  of  witnessing  the  discovery  and  development 
nf  a  still  more  striking  instance  of  this  nature  at  the  Mount  Lyell 
mine  in  Tasmania.  This  is  a  deposit  of  massive  iron  pyrites  about 
300  feet  iu  width,  and  averaging  4^  to  5  per  cent,  copper,  with  3 
pennyweights  gold  and  3  ounces  silver  per  ton  (0.005  percent,  gold 
and  0.01  per  cent,  silver).  At  one  portion  of  the  deposit  there  is  an 
extensive  and  remarkable  chute  of  gossa:i  on  the  footwall,  descend- 
ing to  the  200-foot  level.  This  gossan  contains  nearly  the  same 
amount  of  gold  as  the  original  pyrites,  from  which  it  was  doubtless 
derived,  but  neither  silver  nor  copi^er.  These  commercially  im- 
portant constituents  had  been  leached  out  and  redeposited  on  the 
footwall,  under  the  lower  border  of  the  gossan,  in  a  series  of  exten- 
sive and  irregular  pockets,  which  are  still  being  actively  worked. 
The  first  50  tons  of  ore  extracted  averaged  very  close  to  2,000 
ounces  silver  per  ton  (nearly  7  per  cent.)  and  21  per  cent,  copper. 
The  ore  was  a  pure  chalcopyrite,  containing  streaks  and  nodules 
of  a  very  rich  copper-silver  glance. 

In  our  own  country,  the  United  Verde  mine  in  Arizona  has 
a  layer,  very  rich  in  silver,  between  the  gossan  and  the  ordinary 

Chalcocite,    Copper   Glance,  CuaS ;   79.7   Cu,    20.3   S. 

This  ore  is  seldom  found  in  a  condition  of  perfect  purity,  its 
valuable  component  being  frequently  in  part  replaced  by  iron  and 
other  metals.  Its  copper  percentage  rarely  falls  below  55,  and 
even  at  this  low  standard  the  mineral  retains  its  physical  charac- 
teristics, a  slight  diminution  in  its  luster  being  the  principal  dif- 
ference observable.  When  high  in  copper,  it  greatly  resembles 
the  white  metal  of  the  smelter.  Chalcocite  containing  from  GO  to 
74  per  cent,  of  copper  occurs  in  large  amounts  in  the  noted  Ana- 
conda mine,  Butte,  Montana.  Several  of  the  other  Butte  mines 
carry  the  same  mineral,  although,  as  they  approach   the  western 


bouiularies  of  the  district,  it  gradually  pusses  into  boruite  or  pea- 
cock ore.  It  is  also  an  important  ore  in  Arizona,  occurring  in  large 
quantities  near  Prescott,  as  well  as  in  the  Coronado  and  other 
Clifton  mines.  In  New  Mexico,  it  constitutes  virtually  the  entire 
value  of  the  Nacimiento  and  Oscura  Permian  beds.  It  occurs  fre- 
quently in  Texas  in  the  Grand  Belt  mines,  and  is  the  principal 
ore  of  numerous  narrow  fissures  in  the  Middle  and  Atlantic  States. 
In  the  Orange  Mountains  of  New  Jersey,  examined  by  the  author, 
it  was  found  in  a  species  of  shale,  as  an  ore  of  the  following 

Copper 75.20  ;  Sulphur 17.97 

Iron 4.10      Insoluble 1.10 

Manganese ....   1.13  

Silver  (2.37  ounces) 0  01  99.51 

Gold Trace 

BoRNiTE   OR  Erubescite,  3  CUjS,    FE2S3 ;   55.58   Cu,  16.36  Fe, 

28.0U    8. 

This  is  one  of  the  most  beautiful  of  the  sulphureted  ores  of  cop- 
per, being  characterized  in  its  fresh  condition  by  a  superb  purplish- 
brown  color,  which  soon  changes  on  exposure  to  the  air  into  every 
conceivable  hue,  from  a  golden  yellow  to  the  deepest  indigo,  and 
from  a  brilliant  green  to  a  royal  purple.  The  mode  of  occurrence 
of  this  mineral  and  its  limited  extent  of  distribution  as  regards 
depth  indubitably  stamp  it  as  a  product  of  decomposition,  solution, 
and  re-deposition  of  the  metallic  portion  of  the  vein.  Like  copper 
glance,  this  mineral  is  far  from  uniform  in  its  composition,  varying 
in  richness  from  42  to  nearly  TO  per  cent,  of  copper  without  en- 
tirely losing  its  characteristic  colors. 

It  forms  a  frequent  ore  of  the  Butte  mines  in  their  rich  zone, 
which  lies  between  the  leached-out  surface  zone  and  the  unaltered 
sulphides  below. 

Tetrahedrite,  Gray  Copper  Ore,  Fahlore  (Ci'aS,  FeS,  ZnS. 
AgS,  PbS)4  (SB2S3,  AS2S3);  30.4:0   PER  CENT.  Copper. 

Except  in  those  rare  and  highly  argentiferous  varieties  in  which 
the  copper  is  replaced  to  a  greater  or  less  extent  by  silver,  this  is 
seldom  regarded  in  the  United  States  as  an  ore  of  copper. 

Both  its  scarcity  and  its  obnoxious  components  (arsenic,  anti- 
mony, etc.)  prevent  it  use  as  a  source  of  copper  in  this  country, 
where  the  general  purity  of  our  ores  has  established  such  a  high 

COPPER   AND    ITS    ORES.  13 

standard  for  this  metal.  Only  the  most  favorable  circumstances, 
miueralogical,  metallurgical,  and  commercial,  would  render  the 
working  of  non-argentiferous  fahlores  at  all  practicable.  This 
mineral  occurs  in  small  quantities  in  certain  of  the  Butte  copper 
mines,  rendering  their  product  slightly  inferior  to  that  from  the 
oxidized  ores  of  Arizona  or  the  pure  sulphides  of  Vermont.  This 
slight  disadvantage  is,  however,  far  outweighed  by  their  contents 
in  silver,  which  partly  owes  its  presence  to  this  same  arsenical 
mineral.  From  the  San  Juan  region,  Colorado,  an  argentiferous 
tetrahedrite  adds  a  notable  quantity  to  the  production  of  the 
United  States.  It  appears  principally  as  matte  from  the  lead  fur 
naces,  and  as  black  oxide  from  the  Argo  separating  works. 



The  ores  of  copper  are  widely  distributed  over  the  earth's  sur- 
face, and  may  be  fouud  iu  almost  every  geological  formation. 
The  priucipal  copper  districts  of  North  America  may  be  classed  iu 
three  groups: 

I.  The  Atlantic  coast  beds. 
II.  The  Lake  Superior  deposits. 
III.  The  deposits  of  the  Eocky  Mountains  and  Sierra  Nevadas. 


Throughout  its  whole  extent  in  North  America,  the  Atlantic 
coast  is  bordered  by  a  succession  of  parallel  ranges,  which,  by  their 
general  geological  as  well  as  geographical  analogy,  must  be  classed 
in  the  same  system.  They  form  an  unbroken  chain  from  Florida 
to  Labrador,  and  thence,  continuing  their  same  northeasterly 
direction  along  the  coast  of  that  bleak  country,  dip  beneath  the 
waters  of  Baffin's  Bay,  where  they  are  represented  by  a  series  of 
submarine  peaks,  and,  nourishing  the  gigantic  glacier  system  of 
•western  Greenland,*  terminate,  so  far  as  known,  in  Mount  Edward 
Parry,  north  latitude  82  degrees  40  minutes.  Dr.  T.  Sterry 
Hunt's  admirable  researches  have  given  us  a  very  clear  insight 
into  the  origin,  formation,  and  structure  of  this  immense  range  of 
mountains  within  the  confines  of  the  United  States  and  Canada. 
It  consists  essentially  of  metamorphic  rocks — largely  crystalline 
schists — and  is  metal-bearing  to  a  greater  or  less  degree  throughout 
its  entire  extent,  though  only  in  a  few  places  is  copper  found  iu  i 
sufficiently  concentrated  form  to  justify  auy  attempts  at  extraction 

The  main  copper  mineral  of  importance  in  this  range  is  chalco 
pyrite.  In  the  more  northerly  division,  where  there  has  beei 
extensive  glacial  denudation,  this  reaches  unaltered  almost,  or 
quite,  to  the  grass-roots,  while  from  Virginia  to  Tennessee,  where 

*  See  Dr.  Kane's  Ai'ctic  Expedition  for  soundings  taken  in  Baffin's  Bay;  alsc 
Geology  of  Greenland's  Mountains. 


abrasion  has  not  taken  place,  and  where  oxidation  has  been  assisted 
by  climatic  iufluences,  decomposition  with  subsequent  concentra- 
tion is  found  to  a  considerable  depth.  Tlie  result  of  this  is  usually 
a  zone,  rich  in  an  impure  black  oxide  of  copper  containing  a  certain 
proportion  of  sulphur,  which  sometimes  occurs  in  considerable 
quantities  near  the  surface,  after  first  passing  through  a  greater  or 
less  extent  of  barren  iron  oxide,  derived  from  pyrite,  and  which 
has  no  doubt  furnished  the  copper  to  enrich  the  underlying  zone. 

The  occurrence  of  this  valuable  mineral  in  merchantable  quanti- 
ties has,  in  more  than  one  instance,  raised  expectations  and  led  to 
large  expenditures  that  have  subsequently  proved  entirely  unwar- 
ranted; for  at  a  slightly  greater  depth,  tlie  unaltered  vein  assumes 
its  true  character  of  a  more  or  less  solid  pyrite  or  pyrrhotite  car- 
rying a  very  small  amount  of  copper  (seldom  above  3  per  cent.) 
in  the  form  of  the  common  yellow  sulphuret.  When  the  accom- 
panying mineral  is  a  bisulphide  of  iron  and  the  locality  is  favorable, 
the  pyrite  may  be  utilized  in  the  manufacture  of  sulphuric  aci>l, 
the  copper  being  extracted  from  the  residues  by  well-known 
methods;  but  when  the  prevailing  mineral  is  the  monosu]ph\(]e 
— magnetic  pyrites — there  can  be  no  question  of  profitable  work- 
ing, pyrrhotite  being  absolutely  valueless  since  copperas  has  become 
a  by-product  of  fence-wire  making.  At  Capelton,  in  Canada,  at 
Ely,  Vermont,  and  at  one  or  two  points  in  Newfoundland,  copper 
pyrite  occurs  in  a  sufficiently  concentrated  form  to  yield  from  5  to 
(i  per  cent,  in  considerable  quantities,  an  ore  on  which  profitable 
operations  may  be  conducted,  under  favorable  conditions. 

In  Virginia,  at  Ore  Knob,  North  Carolina,  at  the  Tallapoosa 
mine  in  Georgia,  and  at  Stone  Hill,  Alabama,  indications  of  a 
similar  concentration  of  copper  have  given  rise  to  extensive  explo- 
rations, and,  in  some  cases,  to  the  expenditure  of  large  amounts 
of  money,  which  have  not  always  resulted  satisfactorily.  These 
are  all  examples  of  so-called  bedded  veiiis,  following  the  lines  ot 
stratification,  and  being  simply  sandwiched  in  between  the  layers 
of  rock.  One  of  tlie  most  curious  features  of  these  beds  is  the 
alternate  occurrence  of  the  sulphide  of  iron  that  forms  the  great 
mass  of  the  gangue,  as  pyrrhotite  and  pyrite.  In  Capelton,  for 
instance,  we  have  the  bisulphide;  a  hundred  miles  distant,  at  Ely, 
the  monosulphide  alone  exists;  in  Virginia  and  at  Ore  Knob,  the 
monosulphide  preponderates;  while  in  the  Tallapoosa  mine,  the 
bisulphide  alone  is  found.  Neither  the  chemical  nor  geological 
composition  of  the  corresponding  country-rock  explains  this  phe* 


nomenou.  Here,  it  will  be  proper  to  mention  the  occurrence,  ii, 
stratified  rocks,  of  the  siilpliide  of  copper  (copper  glance),  usually 
in  unimportant  quantity,  throughout  Pennsylvania,  New  Jersey, 
and  other  Middle  and  Southern  States. 

For  convenience,  we  may  append  to  this  division  the  copper  ores 
of  Sudbury,  iu  the  Province  of  Ontario,  Canada.  Copper  pyrites 
is  here  associated  with  much  greater  amounts  of  nickeliferous 
pyrrhotite,  and  occurs  iu  stockwerks  in  the  Huronian  rocks,  along 
or  near  contacts  of  diorite  and  gneiss,  or  diorite  and  quartz-syenite. 


These  deposits  occur  in  the  Keweenawan  series,  which  are  up- 
turned rocks  of  Algonkian  age  that  have  been  deposited  uncon- 
formably  upon  the  iron-bearing  Huronian  series,  and  are  in  turn 
overlapped  by  the  sandstones  of  the  Cambrian. 

The  cojiper-bearing  strata  of  the  Keweenawan  series  consist  of 
beds  of  trap,  sandstone,  and  conglomerates  of  doubtful  age.  They 
rise  at  an  angle  of  4:5  degrees  out  of  the  horizontal  sandstone  from 
which  the  basin  of  Lake  Superior  has  been  eroded.  It  is  only  on 
the  Keweenaw  promontory  of  Michigan  that  they  have  yielded 
copper  in  profitable  amounts,  though  the  same  series  of  rocks, 
always  containing  a  certain  proportion  of  this  metal,  stretches 
westward  across  Wisconsin,  far  into  Minnesota.  In  the  latter 
State,  sulphurets  of  copper  are  sometimes  present  in  the  Keweenaw 
belt,  but  in  Michigan  the  copper  occurs  exclusively  in  the  metallic 
condition,  and  is  believed  to  be  derived  from  the  solutions  formed 
from  the  oxidation  of  the  cupriferous  sulphides  that  abound  iu 
the  underlying  Huronian  formation. 

Three  classes  of  de[)osits  have  been  exploited  on  the  Keweenaw 

1.  Veins  which  iu  some  instt^nces  cut,  and  iu  others  are  parallel 
with  the  beds,  but  which  are  filled  with  vein  stone  different  from 
the  intersected  rocks.  It  is  from  these  veins  that  the  great  masses 
of  native  copper  have  been  derived  that  have  made  such  an  im- 
pression upon  the  public. 

"2.  Copper-bearing  beds  of  amygdaloidal  diabase,  locally  called 
as?i  beds.aud  amygdaloidal  traps. 

3.  Beds  of  conglomerate,  of  which  the  cementing  material  con- 
sists in  part  of  copper.* 

*"The  Copper  Resources  of  the  United  States,"  by  James  Douglas,  Journal 
of  tfie  Society  of .  I  rts,  London. 


The  mass  mines  become  poorer  in  depth,  and  are  considered 
somewhat  hazardous  enterprises. 

The  ash  beds  have  been  much  more  profitable,  the  Quincy, 
Franklin,  Pevvabic,  and  many  other  noted  mines,  belonging  to  this 

The  conglomerate  beds  produced  in  1893,  85,662,000  pounds  of 
fine  copper,  or  75  per  cent,  of  the  entire  output  of  the  Lake  Supe- 
rior District.  Yet  this  large  production  comes  from  a  single  ore- 
chute,  about  three  miles  in  length,  and  penetrated  to  nearly  4,000 
feet  in  depth.  On  this  chute  are  situated  the  Calumet  &  Hecla, 
Tamarack, and  Atlantic  Mines. 

The  average  contents  of  the  ores  now  mined  at  Lake  Superior 
may  be  placed  at  about  2.9  per  cent. 

III. — THE      DEPOSITS      OF     THE      ROCKY      MOUNTAINS      AND      THE 

This  division  includes  a  heterogeneous  collection  of  districts  and 
formations,  and  comprises  nearly  one-half  the  area  of  the  United 

The  rock  formations  of  the  different  mining  regions  in  this  dis- 
trict happen  to  differ  sufliciently  to  enable  us  to  subdivide  it  ac- 
cording to  its  geological  characteristics.*     We  have: 

(-4)  Precambrian  deposits^  which  include  the  celebrated  mines 
of  Butte,  Montana,  and  probably  the  United  Verde  group  of 
Prescott,  Arizona. 

{B)  Palmozoic  deposits^  always  associated  with  eruptive  rocks 
in  the  profitable  mines  of  this  country.  The  most  important 
areas  of  this  class  are: 

(a)  Bisbee,  Clifton,  and  Globe,  in  Arizona,  in  which  the 
ore  occurs  mainly  in  the  lower  carboniferous  lime- 
stones, and  is  ordinarily  oxidized  to  a  depth  of  some 
hundreds  of  feet. 
{b)  Leadville,  Colorado,  the  copper  sulphides  being  found 
in  conjunction  with  pyritous  silver  ores  in  silurian 
limestones  in  contact,and  fracture-planes. 
(c)  Tintic,  Utah,  in  Palaeozoic  limestones,  with  ores  of  gold 
and  silver. 

*  "  The  Geological  Distribution  of  the  Useful  Metals  in  the  United  States." 
See  paper  by  S.  F.  Emmons,  Transactions  American  Institute  Mining  Engineers, 
Vol.  XXII. ,  p.  53,  from  which  is  also  taken  certain  information  regarding  the 
geology  of  the  Lake  Superior  district. 


{C)  Mesozoic  Deposits. — The  most  important  of  these  deposits, 
from  the  commercial  staudpoiut,  are  the  pyritous  beds  of  Cali- 
fornia, that  occur  along  the  foothills  of  the  Sierra  Nevada,  at  the 
contact  of  diabase  and  the  upturned  cretaceous  slates.  In  Texas 
and  Colorado,  and  especiall)'  in  New  Mexico,  there  are  areas  of 
Trias  that  show  a  wide  distribution  of  disseminated  copper,  and 
some  few  points  of  sufficient  concentration  to  warrant  exploitation. 
In  the  cretaceous  throughout  the  Rocky  Mountains,  copper  occurs 
to  a  subordinate  degree,  in  connection  with  ores  of  the  precious 
metals,  whence  there  is  a  considerable  production  of  the  less  valu- 
able metal,  as  a  by-product. 

{D)  Tertiary  and  Recent  Deposits. — No  copper  worth  mention- 
ing is  produced  in  the  United  States  from  such  rocks.  There  are, 
however,  several  spots  in  Arizona  and  New  Mexico,  where  there 
are  recent  deposits  resulting  from  the  surface  leaching  of  copper 
minerals  situated  in  the  older  rocks. 

T/ie  Butte  Mines.* — The  most  northerly,  and  by  far  the  most 
important  of  all  the  ores  included  in  this  divison,  are  the  deposits 
of  Butte,  Montana,  their  output  for  1893  being  155,000,000  pounds 
of  fine  copper,  or  about  69,200  tons  of  2,240  pounds.  This  enor- 
mous production  came  from  a  little  granitic  area  of  (probably) 
precambrian  rocks,  not  over  one  mile  wide  by  two  miles  long,  situ- 
ated on  the  western  slope  of  the  main  divide  of  the  Eocky 

The  ore  occurs  in  irregular  lodes  in  the  granite,  having  an  east 
and  west  strike^  and  an  average  dip  of  some  12  degrees  to  the 
south  from  the  vertical,  though  in  places  this  becomes  as  much  as 
45  degrees.  The  distribution  of  the  ore  is  also  very  irregular, 
extensive  bodies  of  the  same  being  frequently  found  on  breaking 
through  what  appears  to  be  a  well-defined  wall.  Again,  there 
will  be  no  definite  line  between  the  vein  and  the  adjacent  country 
rock.  The  ore  is  usually  found  in  chutes  that  often  extend  for 
several  hundred  feet  along  the  strike,  before  pinching  out.  Their 
depth  is  frequently  even  greater  than  their  length,  though  they 
are  sometimes  broken  by  small  faults. 

*  For  a  detailed  description  of  the  mines  and  metallurgical  works  of  this 
district,  see  a  paper  by  the  author,  entitled  "  The  Mines  and  Reduction  Works 
of  Butte  City,  Montana,"  United  States  Geological  Survey,  3fineral  Resources, 
Albert  Williams,  Jr.,  1885.  For  a  recent  and  more  valuable  description  of  the 
mines  of  Butte,  see  "  The  Ore  Deposits  of  Butte  City,"  by  R.  G.  Brown,  Amer- 
ican Institute  Mining  Engineers,  October,  1894.  1  have  used  this  paper  freely 
in  tlie  present  section. 


The  veins  are  rarely  banded,  and  vary  in  size  from  a  few  inches 
af  compact  ore  up  to  100  feet  or  more,  as  in  the  Anaconda  mine. 
Five  or  six  feet  may  be  regarded  as  the  ordinary  width,  though  a 
large  proportion  of  the  ore  raised  in  Butte  comes  from  stopes  of 
much  greater  width  than  this.  The  gangue  rock  is  usually  gran- 
itic and  silicious,  but  not  quartzose. 

The  croppings  of  the  copper  veins  are  moderately  prominent, 
and  consist  of  the  usual  brownish,  iron-stained  quartz  that  may 
be  found  at  almost  any  point  in  the  great  American  mountain 
chain,  from  Alaska  to  Patagonia.  Just  below  the  surface,  red  and 
yellow  oxides  of  iron  appear,  carrying  high  values  in  silver  and 
gold,  but  usually  low  in  copper. 

These  decomposed  ores  extend  to  the  water  level,  which  is  reached 
at  a  depth  of  from  40  to  300  feet,  depending  upon  the  surface 
irregularities.  At  this  point  begins  the  zone  of  rich,  secondary 
copper  ores  that  have  made  Butte  so  famous.  The  copper  minerals 
of  this  zone  are  difficult  to  determine,  as  they  pass  through  all 
gradations  from  pure  chalcocite  down  to  chalcopyrite,  bornite  being 
also  of  very  frequent  occurrence.  Iron  pyrites  is  usually  present 
in  considerable  amounts. 

Naturally,  these  rich,  secondary  ores  have  fallen  oS  in  depth, 
Ledoux  estimating  their  average  decline  at  2  per  cent,  copper  per 
100  feet.  But  this  diminution  lessens  as  greater  depth  is  gained, 
and  the  ore  raised  from  the  Butte  mines  at  present,  omitting  a 
few  bonanza  bodies,  averages  about  6^  per  cent,  copper  and  5 
ounces  silver  to  the  ton  of  2,000  pounds  (0.017  per  cent,  silver). 
There  is  a  loss  of  some  18  per  cent,  in  concentration,  and  to  this 
must  be  added  the  smelting  loss,  which  will  reduce  the  yield  of 
the  great  bulk  of  the  Butte  ores  to  5  per  cent,  copper  and  4  ounces 
(0.014  per  cent.)  silver. 

Notwithstanding  this  great  decline  in  percentage  and  values 
(which,  to  be  sure,  has  resulted  partly  from  the  ability  to  work 
lower  grade  ores  to  advantage),  the  Butte  mines  are  making  more 
profit  to-day  from  a  6  per  cent,  ore  than  formerly  from  one  of 
double  this  richness.  This  results  from  the  consolidation  of  min- 
ing properties,  and  from  the  astounding  and  radical  improvements 
made  in  the  metallurgical  treatment  of  the  ore.  The  rich  surface 
ores  of  the  district  have  furnished  the  capital  that  was  needed  to 
design  and  construct  the  improved  plants,  and  to  gain  exnerience 
necessary  for  treating  the  lower  grade  ores  at  a  profit. 


Ledonx*  makes  the  following  four  statements,  with  which  1 
agree  in  the  main : 

1.  Tiie  average  yield  of  copper  in  the  Butte  camp  is  5^  per  cent., 
or  110  pounds  per  ton  net,  and  it  will  not  fall  much  below  5  per 

2.  This  copper  costs  9|  cents  per  pound,  delivered  in  New  York. 

3.  The  value  of  the  precious  metals  in  the  copper  is  equal  to 
$57  per  ton  copper,  and  should  yield  a  net  profit  of  over  2  cents 
per  pound  of  copper,  with  silver  at  Go  cents  per  ounce,  and  elec- 
trolytic copper  at  9^  cents  per  pound.  (This  was  written  a  year 

4.  The  present  output  can  be  maintained  for  at  least  ten  years 
to  come. 

At  the  present  low  price  of  the  metals  in  question,  it  may  be 
assumed  that  the  net  profits  of  the  Butte  mines  are  mainly  derived 
from  their  silver  contents,  the  copper  just  about  paying  all  the 
expeuses.  Asiile  fruiu  Lhe  L'uiuuiet  &  Hecla  ore  cliute  at 
Lake  Superior,  this  is  probably  doing  better  than  any  other  great 
copper  district  in  the  world. 

At  a  depth  exceeding  1,300  feet,  there  is  no  sign  of  any  weaken- 
ing or  giving  out  of  the  Butte  copper  lodes. 

The  Arizona  Copper  Mines. \ — These  comprise  four  distinct 
groups  of  deposits  of  commercial  importance,  besides  a  very  large 
number  of  slightly  developed  districts,  some  few  of  which  may  yet 
become  producers.  The  production  in  1893  was  about  44,000,000 
pounds,  or  19,643  tons  of  2,240  jiounds. 

The  profitable  mines  have  been  found  mostly  in  carboniferous 
limestone,  and  at,  or  near,  its  contact  with  an  eruptive  rock,  such 
as  felsite,  diorite,  or  porphyry.  On  entering  an  underlying  acid 
rock,  whether  sandstone  or  porphyry,  the  veins  become  narrow 
and  unprofitable.  Tlie  productiveness  and  permanency  of  most 
ai  the  Arizona  copper  districts  seem  to  stand  in  close  relation  to 
the  thickness  of  the  ore-bearing  limestone.  A  striking  example 
jf  this  fact  may  be  seen  in  the  accompanying  cut,  Fig.  1,  which 
lihows  a  section  across  the  well-known  Longfellow  mine  of  Clifton, 

*  The  Mineral  IndtiMry,  Vo].  II.,  p.  245. 

f  The  cuts  and  many  of  the  facts  in  this  description  are  from  A.  F.  Wendt'.«! 
l^aper,  "  The  Copper-Ores  of  the  Southwest,"  Transactions  American  Institute 
Mfining  Engineers,  Vol .  XV. ,  p.  25. 


The  aciil  rocks,  such  as  diorite,  porphyry,  aud  granite,  contain 
hirge  numbers  of  veins  carrying  copper  ores  with  qnartzose  gangue, 
but  they  have  scarcely  ever  proved  productive  in  this  region.  It 
is  an  interesting  fact  that  in  these  acid  veins,  the  surface  carbon- 
ates and  oxides  usually  change  within  a  few  feet  into  copper  glance, 
and  at  no  very  great  depth,  into  the  ordinary  chalcopyrite,  while 
the  limestone  veins  carry  great  bodies  of  oxidized  ores  to  very  con- 
siderable depths,  and  change  into  chalcopyrite  without  any  marked 
appearance  of  copper  glance. 

As  Wendt  Justly  remarks,  all  the  important  Arizona  deposits 
seem  to  be  true  fissure  veins,  in  the  sense  that  they  are  bodies  or 
masses  of  ore  deposited  in  the  rocks  that  now  contain  them,  subse- 
quent to  the  deposition  or  formation  of  these  rocks. 

Great  bodies  of  clay  are  almost  invariably  found  in  conjunction 
with  these  veins,  resulting,  evidently,  from  the  decomposition  of 
the  rocks  due  to  the  enormous  thermal  action  that  has  taken  place 
during  the  deposition  of  the  copper  ores.  The  walls  of  the  Long- 
fellow mine  often  consist  of  pure  white  kaolin,  of  which  Wendt 
gives  the  following  analysis: 

Silica 42.40 

Alumina 32.50 

Ferric  oxide 16.17 

Lime 2.10 

Magnesia Trace. 

Copper Trace. 

The  balance  of  the  100  per  cent,  was  principally  moisture. 
The  four  important  Arizona  copper  districts  are  at  present: 
The  Clifton  District.  The  Globe  District. 

The  Bisbee  District.  The  Elack  Eange  District. 

It  is  only  possible,  in  this  brief  sketch,  to  outline  a  few  of  the 
most  important  characteristics  of  these  interesting  deposits. 

TJie  Clifton  District^  like  most  of  the  other  copper  areas,  con- 
tains three  distinct  systems  of  veins  carrying  copper. 

1.  Veins  occurring  in  limestone. 

2.  Veins  occurring  in  porphyry  or  felsite. 

3.  Veins  occurring  in  granite. 

The  ores  of  the  first  system  consist  mainly  of  cuprite,  in  a  gangue 
of  compact  hematite:  and  of  malachite  and  azurite,  in  a  gangue  of 
manganese,  or  wad.  An  analysis  of  a  characteristic  specimen  of 
this  cupriferous  wad  by  Professor  Mayer  yielded: 



Cupric   oxide 28.39 

Manganic   oxide 31.24 

Silica 24.81 

Water 11-87 

Ferric  oxide  and  carbonic  acid 2.74 

Lime Trace. 


The  most  noted  mine  of  this  class  is  the  Longfellow.  A  refer- 
ence to  Fig.  1  will  show  it  to  be  an  almost  vertical  fissure  in  strati- 
fied limestone,  at  or  near  its  junction  with  a  strong  dyke  of  felsite. 


Fig.  1. — The  Longfellow  Mine. 

At  times  the  vein  forms  an  actual  contact  with  the  felsite.  Ex- 
tensive bodies  of  ore  branch  from  the  main  vein,  replacing  one  or 
more  beds  of  the  limestone,  and  again  following  vertical  seams  in 
the  latter.  T'igs.  2  and  3  show  horizontal  and  vertical  sections  of 
vein  structure  in  the  Longfellow  mine. 

The  Detroit  mine  also  occurs  in  the  same  carboniferous  lime- 
stone, in  close  proximity  to  a  dyke  of  fine-grained  green  eruptive 
rock.     Its  ores  are  mainly  azurite  and  cupriferous  wad. 

The  second  class  of  veins  occurs  in  porphyry,  and  presents  too 
varied  features  for  detailed  description  in  this  connection.  One 
of  them  is  shown  on  Metcalf  Hill,  see  Fig.  4,  where,  at  the  surface, 
it  forms  a  stock werk  of  oxidized  veinlets  over  100  feet  wide  in  the 
porphyry,  which  soon  unite  into  a  single  vein  carrying  copper 
glance,  at  greater  depths  deteriorating  into  unprofitable  ores. 



Another  interesting  example  of  the  second  system  of  veins   is 
the  Coronado  group,  in  a  strong  dyke  of  quartz  porphyry,  cutting 

Fig.  3. 

Fig.  2.— Horizontal  Section.  Fig.  3.— Veutical  Section. 

The  Longfellow  Mine. 

•X  X  XX 


Fig.  6. — Bisbee  Deposits. 

through  syenite  and  granite,  which  latter  abuts  against,  and  is  sur- 
rounded by,  stratified  limestone,  as  shown  in  Fig.  5.  Near  the 
surface,    these    veins   carry  strong  bodies  of   rich  copper  glance, 



mainly  where  the  poiphyritic  walls  are  strongly  decomposed  and 
kaolinized.  As  depth  is  gained  the  rich  ore  gradually  disappears, 
and  at  150  to  :200  feet  from  the  surface  the  vein  becomes  barren, 
or  contains  only  sparsely  disseminated  chalcopyrite. 

Two  partial  analyses,  by  Henrich,  of  typical  ores  of  this  class 
show  their  silicions  character: 

I.  II. 

Copper 11. IT  21.95 

Silica... 67.00  48.90 

Iron 6.91  9.41 


*..  \PORPHYR(Y      ++  * 
+     +  *  v+  lit  ■*.  +     *  +•* 

SKCTION"    of    METC.VI.F    IIll.I.,   Cl.IKTOX. 

The  veins  of  the  tliird  system  occur  in  granite,  at  a  great  alti- 
tude and  in  extremely  inaccessible  situations.  They  are  strong 
and  of  good  width — 5  to  I'i  feet — and  carry  copper  glance  near  the 
surface;  but  their  mineralization  is  very  irregular  and  they  have 
been  little  worked. 

77/f  Bisbee  Disfn'rf  is  in  the  Mule  Pass  Mountains,  in  Southern 
Arizona,  only  10  miles  from  the  Mexican  border. 

A  great  mass  of  eruptive  rock  has  upheaved  the  carboniferous 
limestone,  and  along  the  southern  contact  occurs  the  Copper  Queen 
group  of  deposits.  (See  Fig.  0.)  They  correspond  closely  to  Von 
Cotta's" bed-veins."  They  are  not  simply  "ore-beds,"  as  they  send 
numerous  spurs  into  the  walls.  These  spurs  usually  follow  the 
planes  of  bedding  of  the  limestone,  and  the  mode  of  deposition  is 


Still  farther  complicated  and  obscured  by  the  ocenrreuce  of  affiliated 
bodies  of  ore  in  the  limestone,  which  were  evidently  deposited  in 
vugs  and  caves. 

The  ore  consists  mainly  of  hydrated  oxides  of  iron  and  alnraiua, 
carrying,  at  present,  about  8  per  cent,  of  copper,  after  undergoing 
a 'moderate  selection.  To  a  depth  of  over  400  feet  the  copper 
was  mainly  in  the  form  of  carbonates,  but  as  greater  depth  is 
gained  sulphides  are  encountered,  and  a  converter  plant  has  just 
been  erected. 

The  Bisbee  black  copper,  as  produced  by  a  single  fusion  of  the 
oxidized  ores  with  coke,  in  a  water-jacket  cupola,  is  of  excellent 






X^xHj,'    X 

^""mx  x" 





Fig.  5. — Vertical  Crosssectiox  of  Cohoxado  Vein,  Clifton. 

quality,  the  following  analysis  by  the  Orford  Copper  Company, 

representing  one  lot  of  60  tons. 

Oopper 95.00 

Sulphur 0.44 

Iron 4.23 

Insoluble 0.51 

Arsenic None. 

Antimony , None. 


The  Globe  Diftfrict  is  situated  more  toward  the  center  of  Ari- 
zona, on  the  eastern  slope  of  the  Pinal  Mountains. 

The  main  ore  body  that  has  made  this  district  famous,  is  situ- 
ated, as  usual,  in  carboniferous  limestone,  close  to  an  upheaval  of 
diorite.     (See  Fig.  7.) 


An  analysis  by  Dr.  Trippel,  of  a  week's  delivery  of  ore  to  the 
furnaces,  gives: 

Silica 20.23 

Ferric  oxide 42. 10 

Alauiitia 4.15 

Loss  by  ignition 9.75 

Oxide  of  copper 17.12 

Magnesia , 2.85 

Lime 1.12 

Oxide  of  manganese 1.63 


Fig.  7.  — Skctiox  of  Globe  Mixe. 

This  sample  is  more  ferruginous  than  the  general  run  of  the 
ore,  which  usually  requires  the  addition  of  limestone  before 

Very  pure  black  copper  is  produced  by  a  single  fusion  with  coke, 
in  water-jacket  furnaces,  the  following  analysis  by  Trippel  being 
a  sample  of  two  weeks'  production,  which  is,  however,  slightly 
above  the  average  in  purity: 


Copper t 99,11 

Lead 0.67 

Sulpbur 0.08 

Slag 0.08 

Arsenic  Trace. 

Iron Trace. 


The  Blach  Range  Copper  District  is  situated  near  the  center  of 
Arizona,  on  the  eastern  slope  of  the  Black  Kange,  and  close  to  the 
Verde  river. 

The  veins  occur  near  the  contact  of  a  belt  of  diorite  and  slate, 
and  are  of  great  strength.  The  non-argeutiferons  green  carbon- 
ates and  oxides  give  way  to  massive  pyrite  and  chalcopyrite  at  a 
depth  of  about  150  feet.  These  sulphides  contain  moderate 
amounts  of  silver  and  gold,  the  oxysulphureted  ore  (often  called 
black  oxide),  at  the  junction  of  the  oxidized  and  sulphide  ores, 
being  often  extremely  rich  in  the  precious  metals.  Very  extensive 
and  valuable  bodies  of  pyritic  ores  have  been  lately  developed,  and 
are  being  smelted  and  converted  on  the  ground,  and  a  railroad  is 
building  to  the  mine. 

The  mining  districts  of  Lake  Superior,  Butte,  and  Arizona  fur- 
nish about  95  per  cent,  of  the  total  copper  produced  in  the  United 



The  first  step  usually  taken  in  the  treatment  of  an  ore  of  copper 
is  to  learn  its  value  by  determining  the  proi)ortion  of  that  metal 
that  it  contains.  This  process  is  called  assaying,  as  distinguished 
from  chemical  analysis,  which  includes  the  further  investigation 
as  to  the  general  composition  of  the  ore. 

We  shall  confine  our  discussion  in  this  place  to  assaying  only. 
The  assaying  of  any  given  parcel  of  ore  is  necessarily  preceded  by 
the  process  of  sampling^  by  which  we  seek  to  obtain,  within  the 
compass  of  a  few  ounces,  a  correct  representative  of  the  entire 
quantity  of  ore,  which  may  vary  in  amount  from  a  few  pounds  to 
several  thousand  tons.  With  rich  ores,  it  will  lessen  the  chance 
of  serious  error  in  large  transactions  to  divide  the  lot  into  parcels 
of  not  over  fifty  tons  each,  and  sample  each  of  these  lots  by  itself. 

The  utmost  care  and  vigilance  in  sampling  and  assaying  should 
be  required  at  every  smelting  works,  both  in  the  interest  of  the 
works  and  in  that  of  tlie  ore-seller. 

American  conditions  have  encouraged  the  use  of  automatic 
devices  for  the  sampling  of  ores  and  mattes,  and  although  there  is 
still  a  certain  prejudice  against  them  in  some  private  works,  I  be- 
lieve that  they  have  been  adopted  by  all  public  sampling  works  of 
any  standing. 

Such  works  are  constantly  handling  large  quantities  of  rich  and 
very  varied  material,  and  it  is  a  matter  of  absolute  necessity  to 
them  that  their  methods  of  sampling  should  be  above  suspicion, 
and  free  from  the  factor  of  **  personal  equation"  that  would  be 
introduced  by  the  employment  of  a  reasoning  agent  to  take  the 

Automatic  samplers,  constructed  on  correct  principles,  must 
necessarily  attain  absolute  accuracy,  and  a  sufficiently  extended 
comparison  of  their  resnlts  with  those  obtained  by  hand-sampling, 
will  satisfy  any  one  of  their  superiority. 



The  methods  of  haud-sampliiig  are  too  well  known  to  demand 
description  in  these  pages. 

Aatomatic  samplers  may  be  divided  into  two  classes: 

1.  Those  which  divert  a  portion  of  the  falling  stream  of  ore, 
either  constantly  or  intermittently. 

2.  Those  that  divert  the  entire  ore-stream,  for  an  instant,  at 
regular  intervals  of  time.* 

■fl     o- 



1  i^b(i  r^- 

Fig.  8.— Brunton's  Sampler. 

The  devices  of  the  first  type  are  very  numerous.  Some  of  them 
are:  A  cone,  or  dividing-box,  upon  which  the  crusher  discharges, 
and  which  automatically  separates  from  one-third  to  one-tenth  of 
the  whole.  The  sample  thus  obtained  can  be  still  further  dimin- 
ished by  successive  operations  on  similar,  but  smaller  apparatus, 
lower  down.  Or,  a  wedge  is  used  to  separate  the  falling  ore- 
stream  into  a  very  large,  and  a  very  small  portion. 

*The  tables,  and  much  of  the  text  that  follows,  are  adapted  from 
Dr.  Ledoux's  paper  ou  "  American  Methods  of  Sampling  and  Assaying  Copper," 
The  Mineral  Industry,  Vol.  1. 


Many  ingenions  machines  exist  for  accomplishing  the  same  end 
by  various  means;  but  none  of  them  have  been  entirely  satisfac- 
tory, owing  to  the  tendency  of  the  coarse  and  fine  particles  of  ore 
to  segregate,  and  thus  to  render  the  ore-stream  richer  laterally,  or 
in  the  center.  And  on  different  ores  the  relative  position  of  these 
rich  and  poor  streaks  may  vary  completely. 

Hence,  we  must  turn  to  the  second  class  of  automatic  samplers 
— those  that  momentarily  divert  the  entire  falling  ore-stream  for 
a  sample.  On  well-planned  machines  of  this  description,  foreign 
substances,  such  as  rags,  chips,  frozen  lumps  of  ore,  etc.,  produce 
no  effect  inimical  to  accuracy. 

brunton's  automatic  sampler. 

This  machine  deflects  the  entire  ore-stream  to  the  right  or  left, 
while  falling  through  a  vertical  or  Inclined  spout.     By  a  simple 

Fig.  9. — Brunton's  Quartering  Shovel. 

arrangement  of  movable  pegs,  in  connection  with  the  driving  gear, 
the  proportion  of  the  ore-stream  thus  deflected  into  the  sample-bin 
may  vary  from  10  to  50  per  cent.;  the  latter  amount  only  being 
required  in  coarse  ores  of  enormous  and  very  variable  richness, 
while  for  ordinary  lump  ores,  from  10  to  20  per  cent,  is  the 
maximum  required. 

Instead  of  passing  the  sample-stream  of  ore  into  a  bin,  this 
system  may  be  still  further  perfected  by  leading  it  directly  to  a 
pair  of  moderately  fine  rolls,  the  product  of  which  is  elevated  to  a 
second  similar  sampling  machine,  from  which  the  final  sample 
drops  into  a  locked  bin,  to  be  pulverized  and  quartered  by  hand. 

The  two  macliines  are  driven  at  different  speeds,  to  prevent  any 
possible  error  that  might  rise  from  isochronal  motion. 

A  still  more  recent  invention  of  Mr.  Brunton's  is  the  quarter- 
ing shovel,  described  in  the  Engineering  and  Mining  Journal  of 
June,  1891. 

H.   L.   Bridgmau   has  invented   and    introduced   an   automatic 


sampler  whose  principle  is  so  sound  and  results  attained  so  satis- 
factory, that  I  feel  obliged  to  describe  it  at  some  length,  I  make 
use  of  a  portion  of  Mr.  Bridgraan's  description  and  illustrations.* 

Machine  A. 

This  machine  occupies  a  floor-space  of  3  by  4  feet,  and  has  a 
total  height  of  7  feet  6  inches.  It  is  self-contained,  requiring 
only  to  be  bolted  to  the  floor  and  to  have  feed,  discharge,  and 
belt  connections  made.  Fig.  10  shows  the  machine  as  it  is  built, 
while  Figs.  11  and  12  give  the  diagraphic  sections  and  details, 
some  minor  changes  and  omissions  having  been  made  for  the  sake 
of  clearness.  The  machine  consists  essentially  of  three  appor- 
tioners,  I,  II,  and  III,  all  driven  by  the  one  pulley,  X  (usually 
tight  and  loose  pulleys),  and  three  stationary,  concentric  recepta- 
cles. El,  Kg,  and  H,  so  constructed  that  any  material  falling  into 
them  will  pass  out  through  the  spouts  T^  and  Tg  into  the  sample- 
buckets  Z^  and  Za  or  through  the  spout  S,  which  discharges  the 
rejected  portion  of  the  sample.  Apportioners  I  and  III  revolve 
in  the  same  direction,  apportioner  II  in  the  opposite  direction;  I 
at  about  5,  II  at  about  15,  and  III  at  about  45  revolutions  a  min- 
ute. That  is  to  say,  each  apportioner  moves  actually  three  times 
as  fast  as  the  one  above  it,  and  in  the  contrary  direction,  or,  rela- 
tively, four  times  as  fast.  By  the  use  of  this  expedient  of  contrary 
revolution,  the  same  relative  speeds  are  obtained  as  though,  all 
revolving  in  the  same  direction,  the  actual  speeds  were  respectively 
5,  25,  and  125,  at  which  latter  speed  centrifugal  force  would 
become  very  troublesome. 

The  upper  apportioner,  I,  consists  of  two  concentric  rings, 
divided  by  8  partitions  into  8  equal  topless  and  bottomless  com- 
partments, L,  from  each  one  of  which  leads  an  adjustable  spout, 
either  as  M-1,  or  as  M-2,  or  as  M-D.  Set  in  rotation,  spout  M-1 
would  describe  a  certain  circular  path,  1-1;  spout  M-2  a  certain 
other  path,  2-2,  and  spout  M-D  a  third  path,  W  (see  Fig.  12). 

The  intermediate  apportioner,  II,  is  merely  a  conical  funnel, 
having,  besides  the  large  outlet  W,  four  vertical  shoots,  IS'i-Nj 
and  N2-N2,  through  its  sloping  sides  as  shown  in  Fig.  12;  each 
one  of  these  shoots  forms  one-eighth  of  the  circular  paths  covered 
by  the  spouts  M-1  and  M-2  respectively. 

The  lower  apportioner  III  is  of  the  same  construction  as  II  and 
bears  the  same  relation  to  it  that  II  bears  to  I. 

^IVansactions  American  Institute  Mining  Engineers,  Vol.  XX.,  p.  416. 

Fig.  10. 
Mechanical  Ore  Sampler.     Macliine  A.     (Jeneral  View. 

Fm.  11. 
Mechanical  Ore  Sampler.     Size  A.     Total  Height,  including  Sample  Buckets. 

7  feet  6  inches. 

Fig.  12. 
Mechanical  Ore  Sampler.     Size  A. 


All  example  will  best  illustrate  the  operation  of  the  niacbiue. 
it  may  be  assumed  that  au  original  sample  of  40,960  pounds  (tl.e 
960  being  added  to  avoid  fractious)  is  to  be  put  through  the 
raachiue;  tliat  the  time  required  will  be  one  honr;  that  the  speed 
of  the  machine  is  sucii  that  the  upper  apportioner,  I,  will  make 
320  revolutions  in  that  time,  and  finally  that  the  ore  is  of  such 
grade  and  character  as  to  only  require  the  smallest  sam.ple  that 
the  machine  will  give.  Under  these  conditions,  one  of  the 
spouts,  Ml,  would  be  set  as  M-1,  one  (the  opposite  one)  as  M-2,  and 
the  remaining  six  as  M-D  (Fig.  11). 

The  flow  of  material,  previously  crushed  to  below  one  inch  in 
size,  would  then  be  started  through  the  feed-spout,  F,  and  the 
machine  set  in  motion. 

It  is  evident  that  at  each  revolution  one  320th  part  of  the  whole 
lot,  or  128  pounds,  will  pass  through  the  feed-spout  F.  Of  this 
amount  six-eighths,  or  96  pounds,  will  be  discarded  by  the  six 
spouts,  M-D,  passing  down  through  W,  W,  H,  and  so  through 
the  spout  S  and  out  of  the  machine,  while  one-eighth  of  the  128 
pounds,  or  16  pounds,  forming  the  first  cut  of  the  first  or  outer 
sample,  will  pass  through  the  spout  M-1,  and  the  remaining  one- 
eighth,  or  16  pounds,  forming  the  first  cut  of  the  second  or  inner 
sample,  through  the  spont,  M-2. 

These  two  first  cuts  will  proceed  side  by  side,  by  separate  paths, 
through  the  same  series  of  operations,  and  whatever  applies  to  the 
one,  applies  eqnally  to  the  other;  it  will,  therefore,  suffice  to  fol- 
low the  first  sample.  This  one-eighth,  or  16  pounds,  having  been 
cut  from  the  mass  by  the  partitions  of  the  compartment  Li,  of 
which  M-1  forms  an  extension,  will  drop  nearly  vertically  through 
M-1  on  its  way  to  the  sample  box,  Zj.  As  it  leaves  the  spout, 
M-1,  duriug  the  one-eighth  of  a  revolution  that  is  occupied  by  the 
said  M-1  in  passing  beneath  the  feed-spout,  F,  it  will  be  inter- 
cepted by  the  intermediate  apportioner,  II,  which  in  the  same 
time  will  have  made  3ne  half-revolution  (relatively  to  1). 

Since  the  vertical  shoot,  N-1,  occupies  one-fourth  of  the  semi- 
circumference  of  II  passing  beneath  the  spout,  M-1,  it  follows  that 
one-quarter  of  the  16  pounds,  or  4  pounds,  will  drop  vertically 
through  this  shoot  as  the  second  cut  of  the  first  sample.  The 
remaining  three-quarters,  or  12  pounds,  will  pass  down  the  sloping 
sides  of  II  and  be  discarded  through  W,  W,  11,  and  S. 

In  precisely  the  same  way,  the  second- cut  of  4  pounds  will  b'e 
quartered  by  the  lower  apportioner.  III,  3  pounds  being  discarded 

36  MOUEKN    COl'l'Kli    SMLLTlNtx. 

aud  1  pouui],  as  the  third  aud  final  ctit,  passing  through  the  ver- 
tical shoot  Pi  aud  the  spout  Ti  into  the  sample  bucket  Zj. 

In  the  same  way  a  1-pound  portion,  as  the  third  cut  of  the  sec- 
ond sample,  will  find  its  way  to  the  bucket  Z.,. 

This  series  of  operations  will  occur  at  each  revolution  of  the 
upper  apportioner;  and  at  the  end  of  the  hour  each  of  the  buckets 
Zi  and  Z.,  will  contain  320  portions  of  1  pound  each,  or  a  total 
final  sample  of  'SW  pounds,  these  two  total  samples  being  as  inde- 
pendent of  each  other  as  though  made  at  different  times  aud 
places.  It  will  of  course  rarely  happen  that  this  theoretical  exact- 
ness of  weights  will  obtain,  which  point  will  be  considered  later. 
Should  the  ore  be  of  higher  grade  or  more  irregular  in  character, 
two  or  three  or  four  of  the  spouts,  M,  may  be  set  for  each  sample, 
giving  final  samples  of  040,  060  and  1,280  pounds  respectively. 

It  will  be  noticed  that  only  the  discarded  part  of  the  sample  is 
touched  by  the  machine,  the  retained  portions  dropping  nearly 
vertically  and  practically  freely  through  the  machine,  until,  in  a 
finished  condition,  they  reach  the  stationary  receptacles  Ej  and 
R21  or  the  sample-buckets  Zj  and  Z,.  The  machine  can  have  had, 
therefore,  no  influence  on  the  constitution  of  the  samples,  and 
"coarse"  aud  "'tine"  must  be  contained  therein  in  the  same  pro- 
portion as  delivered  by  the  feed-spout  F. 

It  may  be  remarked  in  passing  that  the  finer  the  material  the 
slower  the  feed,  aud  the  greater  the  speed  of  the  machine  the 
greater  will  be  the  distribution  and,  presumably,  the  better  the 
samples.  The  conditions  above  given,  however,  are  easily  attained, 
depending  only  on  the  crushing  capacity  at  disposal,  and  have 
been  found  by  experience  to  give  satisfactory  results,  it  being  par- 
ticularly desirable  not  to  use  a  much  higher  speed.  For  light  or 
wet  ores  it  may  be  neeccssary,  in  order  to  avoid  an  accumulation 
of  material  in  the  machine,  to  retluce  the  speed  to  half  tliat  given. 
This  lower  speed  may  of  course  be  used  for  heavy  materials  also, 
the  only  practical  difference  between  the  higher  and  lower  speeds 
(aside  from  the  influence  of  centrifugal  force)  being  the  difference 
in  the  number  of  cuts  maile  by  the  machine. 

In  lump  ores,  it  is  difticnlt  to  obtain  a  correct  sample,  even  for 
moisture,  without  some  preliminary  crushing,  and  to  save  labor  it 
is  best  to  use  a  portion  of  the  large  sample  from  the  automatic 
sampler  for  this  purpose;  the  accurate  weighing  of  the  entire  ore 
parcel  being  postponed  until  jnst  before,  or  after,  the  sampling,  and 
the  portion  reserved  for  the  moisture  determination  being  placed 


in  an  open  tin  vessel,  contained  in  a  covered  metal  case,  having  an 
inch  or  two  of  water  on  its  bottom,  in  which  the  sample  tins 

From  one-fonrth  to  one-half  pound  of  the  sample  is  usually 
weighed  out  for  this  determination,  and  dried  under  frequent  stir- 
ring, and  at  a  temperature  not  exceeding  212  degrees.  While  it 
is  always  important  to  keep  within  the  limit  of  temperature  Just 
mentioned,  it  is  especially  the  case  with  certain  substances  which 
oxidize  easily.  Among  these  are  finely  divided  sulphides,  and 
above  all,  the  pulveruleiit  copper  cements  obtained  from  precipi- 
tating copper  with  metallic  iron  from  a  sulphate  solution. 

Such  a  sample,  containing  actually  5-j  per  cent,  of  moisture, 
showed  an  increase  of  weight  of  some  2  per  cent,  on  being  exposed 
for  thirty  minutes  to  a  temperature  of  about  235  degrees  Fahr. 

Certain  samples  of  ore — especially  from  the  roasting  furnace — 
are  quite  hygroscopic,  and  attract  water  rapidly  after  drying. 

In  such  cases,  the  precautions  used  in  analytical  work  must  be 
employed,  and  the  covered  sample  weighed  rapidly,  in  an  atmos- 
phere kept  dry  by  the  use  of  strong  sulphuric  acid. 

The  sampling  of  the  malleable  products  of  smelting,  such  as 
blister  copper,  metallic  bottoms,  ingots,  etc.,  can  only  be  satisfac- 
torily effected  by  boring  a  hole  through  a  certain  proportional 
number  of  the  pieces  to  be  sampled. 

Where  such  work  is  only  exceptional,  an  ordinary  ratchet  hand- 
drill  will  answer,  but  in  most  cases,  a  half-inch  drill  run  by 
machinery  is  employed. 

The  chips  and  drillings  are  still  further  subdivided  by  scissors, 
and  as  even  then  it  is  ditKcult  to  obtain  an  absolutely  perfect  mix- 
ture, it  is  best  to  weigh  out  and  dissolve  a  much  larger  amount  than 
is  usually  taken  for  assay,  taking  a  certain  proportion  of  the  thor- 
oughly mixed  solution  for  the  final  determination. 

Many  of  the  smelters  are  too  careless  in  the  sampling  of  their 
metallic  products.  At  the  public  sampling  works  in  New  York, 
where  much  copper  in  metallic  form  has  been  shipped  abroad  for 
refining  and  separation,  the  following  precautions  have  been  found 
necessary  to  ensure  uniform  results. 

With  copper  bars  that  are  tolerably  uniform,  and  free  from  pre- 
cious metals,  every  fifth  bar  is  bored  halfway  through,  on  opposite 

In  sampling  argentiferous  bars,  every  bar  is  usually  bored  twice. 
If  the  bars  carry  appreciable  quantities  of  gold  (and  always  in  the 



case  of  anodes)  the  borings  are  melted  and  granulated,  or  recast 
into  a  sample  bar,  which  is  again  bored. 

In  sampling  and  assaying  matte  for  shipment,  it  mnst  be  remem- 
bered that  in  the  long  journey  from  the  West  there  is  always  a 
certain  loss  in  weight.  This  is  especially  the  case  wlien  the  matte 
is  shipped  in  pigs  (in  bulk),  and  there  are  one  or  more  transfers. 
The  pigs  grind  against  each  other,  producing  a  considerable  amount 
of  powder,  while  the  brittle  edges  and  corners  are  badly  chipped. 
In  the  hurry  of  transfer  it  is  almost  impossible  to  have  the  cars 
that  are  emptied  cleanly  swept,  and  a  loss  always  occurs,  which,  in 
former  years,  I  have  been  inclined  to  put  at  1  per  cent.* 

Under  somewhat  similar  conditions,  Ledoux  found  a  loss  of  0.8 
per  cent,  on  a  lot  of  about  500  tons  of  matte,  shipped  from  the 
West  in  bulk  and  transferred  once. 

Matte  crushed  and  sacked  may  undergo  a  slight  shrinkage  from 
sifting,  or  a  still  more  serious  one  from  torn  sacks,  if  hooks  are 
used  in  handling  it.  Ledoux  gives  the  following  table  as  a  good 
average  result  where  care  in  sampling  and  sacking  is  used  at  the 
smelter  and  the  material  is  crushed  and  sacked.  It  represents 
various  monthly  shipments  of  matte  from  a  Western  smelter  to  a 
public  sampler  in  Ne\v  York,  and  shows  the  weights  and  assays  at 
each  end  of  the  line. 

Mine  Weight. 

Final  Weight. 


Mine  Assay. 

Final    Assay. 



Per  cent. 

Per  cent. 

Per  cent. 









































Average. . 




There  is  a  difference  of  0.07  per  cent,  against  the  mine  in  the 
weights,  and  of  0.03  per  cent,  in  favor  of  the  mine  in  the  assays. 

*  In  one  instance  a  carload  of  matte  weighed  40,000  pounds.  The  matte 
contained  60  per  cent,  copper,  worth,  at  that  time,  10  cents  per  pound,  and  no- 
precious  metals  worth  separating.  A  loss  of  1  percent.,  therefore,  means  a 
money  loss  of  $24  per  carload,  which,  under  the  conditions  referred  to,  would 
not  pay  the  cost  of  sacking. 



Some  of  the  matte  giveu  in  the  last  table  contained  silver.  Tlie 
following  statement  shows  the  results  of  the  determinations  of  this 

Mine  Assay.  Ounces. 

Final  Assay.  Ounces. 

Difference.  Ounces. 

Average  Difference. 




(Against  the  Mine) 

0.35  Ounces 

Per  Ton. 

On  88  carloads  of  matte  and  bars  shipped  by  the  Pennsylvania 
Salt  Manufacturing  Com.pany  to  a  New  York  sampling  works,  the 
average  total  discrepancy  was  0.03  percent,  copper,  and  0.08  ounces 
silver  per  ton  of  2,000  pounds. 

It  has  been  a  matter  of  great  Importance  to  smelters  and  miners 
in  this  country  to  learn  the  exact  system  of  weighing  and  sam- 
pling practiced  in  England,  in  order  that  they  might  obtain  some 
clue  to  the  heavy  discounts  they  are  often  obliged  to  bear,  both  in 
weights  and  assays.  The  exportation  of  copper  mattes  or  other 
similar  products  from  the  United  States  to  England,  for  refining, 
has  pretty  much  ceased,  as  they  can  be  treated  more  profitably  at 
home;  but  as  there  are  other  countries  which  will  doubtless  continue 
shipping  co|)per-bearing  material  to  Swansea  for  a  long  time  to 
come,  it  may  be  useful  to  describe  the  difEerence  between  the  Eng- 
lish method  of  weighing  and  assaying,  and  our  own.  I  again 
quote  from  Dr.  Ledonx: 

"In  the  United  States,  the  public  samplers — at  least  those  in 
the  East — employ  "sworn  weighers,"  who  have  gone  before  a 
notary  public  and  taken  an  oath  of  office.  Ore  in  bulk  is  weighed 
on  platform  scales  in  barrows,  or  small  wagons,  holding  500  to 
1,500  pounds.  The  weight  is  taken  on  a  rising  beam,  which 
amounts  to  an  allowance  of  one-eighth  to  one-fourth  pound  per 
load.  Ore  or  matte  in  bags  is  always,  where  practicable,  weighed 
on  beam  scales,  six  to  ten  bags  being  taken  for  a  draft.  No 
returns  are  ever  based  on  carload  weights  on  track  scales. 

"The  settlement  is  always  based  on  the  acutal  weight  ascertained 
as  above,  with  no  allowance  for  draftage,  etc.,  and  upon  the  actual 
percentage  of  copper  found  by  analysis,  less  the  arbitrary  deduc- 
tion of  1.3  per  cent,  of  fine  copper.  This  arbitrary  deduction  is 
the  result  of  agreement  be<-weeii  the  co])per  smelters  and  producers 

40  JIODKHN    <()I'l'!:ii    S.MKLTING. 

of  the  United  States,  and  is  supposed  to  represent  the  average  loss 
in  smelting.  In  refining  bars,  there  is,  of  course,  no  such  loss  as  1.3 
units,  and  the  smelter  is  the  gainer;  while  in  leady  mattes  or  base 
ores  there  may  be  a  considerably  greater  loss  than  1.3  per  cent. 
In  America,  the  smelter  protects  himself  in  the  price  he  bids  or 
in  the  reiining  toll  he  charges,  instead  of  asking  the  assayer  to  find 
out  for  him,  in  each  case,  what  his  loss  is  likely  to  be — which  is 
what  the  Cornish  assay  attempts  to  do. 

"I  am  indebted  to  the  Liverpool  Wharf  Company  for  the  fol- 
lowing table,  representing  its  experience  with  some  2,500  tons  of 
matte  from  America: 


Shipping  Weight.  Landing  Weight.  Loss. 

Tons.  Tons.  Per  cent. 

l,355iim.  1.354HIJ.  0.3 


Shipping  Weight.  Landing  Weight.  Loss. 

Tons.  Tons.  Per  cent. 

1,139^^^0.  l,130?m.  0.75 

"Difference  between  American  assay,  after  deduction  of  1.3  per 
cent,  and  Cornish  assay,  1.79  per  cent,  copper,  more  or  less. 

"  Difference  between  American  and  English  assay  for  silver,  0.30 
ounces  per  ton  of  2,240  pounds.* 

*  Occasional  shippers  are  sometimes  embarrassed  by  the  unfamiliar  weights 
and  money  used  in  the  English  shipping  returns. 

The  English  employ  the  long  tim  of  2.2-iO  pounds,  which  they  divide  into 
20  hundredweight,  each  hundredweight  containing  4  quarters  of  28  pounds 

When  these  weights  are  to  be  multiplied  by  a  specified  number  of  pounds, 
shillings,  pence. and  farthings,  it  forms  an  exceedingly  fascinating  problem  for 
an  idle  day;  the  very  uncertainty  of  the  result  adding,  in  no  small  degree,  to 
the  interest  of  the  operation. 

I  append  a  recent  example  of  a  shipment  of  some  high-grade  material  to 
London,  giving  the  final  results  in  both  English  and  American  weights  and 

The  English  returns  gave  29  tons,  17  hundredweight,  3  quarters,  27  pounds, 
at  £217  4s.  3Ad.  per  ton. 

The  translation  into  American  gives  33,487  tons  (of  2,000  pounds)  at  $1,051.32 
equals  $35,205.55.  (Thirty-three  and  four-hundred  eighty-seven-thousandths 
tons  at  ten  hundred  fifty-one  dollars  and  thirty-two  cents,  makes  thirty-five 
thousand  two  hundred  and  five  dollars  and  fifty-five  cents.) 

Those  interested  in  the  higher  mathematics  may  enjoy  calculating  the 
English  returns. 


"In  my  opinioD,  an  average  dednctiou  of  fonrpence  jier  unit  of 
copper  made  by  British  buyers  purchasing  in  the  United  States, 
free  on  board  in  New  York,  and  selling  again  at  English  terms, 
will  be  sufficient  to  cover  the  difference  caused  by  loss  in  transitu 
and  by  the  employment  of  the  Cornish  essay. 

"Considerable  of  the  loss  in  weight  can  be  avoided  if  matte  is 
shipped  in  barrels  instead  of  bags.  Good  kerosene  barrels  can  be 
purchased  at  about  85  cents  apiece.  Glucose  barrels  are  too 


American  assayers  and  chemistsf  are  accustomed  to  exercise 
entire  freedom  in  their  selection  of  method  employed  for  the  de- 
termination of  the  constituents  of  any  material  submitted  to  them. 
It  is  only  required  of  them  that  their  results  be  correct.  Conse- 
quently, they  do  not  make  use  of  the  Cornish  fire-asaay,  which  is 
not  properly  a  method  for  the  determination  of  the  exact  amount 
of  copper  contained  in  an  ore,  but  rather  an  ingenious  adaptation 
of  metallurgical  processes  to  laboratory  conditions,  and  which  is 
intended  to  show  the  amount  of  copper  the  smelter  may  expect  to 
produce  from  the  ore  in  question. 

On  the  whole,  it  is  decidedly  favorable  to  the  smelter;  as  on 
any  ordinary  sulpliide  material  of  tolerable  richness,  it  rarely  gives 
so  high  a  result  as  the  analytical  assay,  less  our  1.3  per  cent,  arbi- 
trary deduction.  And  as  1.3  nnits  has  been  found,  by  long  expe- 
rience, to  be  a  sufficient  deduction  to  cover  the  actual  metallurgical 
loss  in  ordinary  furnace  material,  the  inference  is  obvious.  More- 
over, it  introduces  an  unfortunate  element  of  uncertainty  into  all 
transactions  between  miner  and  smelter,  as  the  different  chemists 
seldom  agree  exactly  in  this  assay,  and  frequently  differ  widely. 
With  us,  a  variation  of  0.2  per  cent,  is  sufficient  to  call  for 
adjustment.  " 

In  assaying  slags  for  copper  it  should  be  borne  in  mind  that  even 
after  apparent  complete  decomposition  by  acid,  theinsoluble  residue 

*  Not  feeling  myself  competent  to  treat  exhaustively  of  the  improvements 
made  in  copper  assaying  during  the  past  fifteen  years,  I  have  availed  myself 
of  the  kind  assistance  of  several  well-known  chemists.  Their  names  will 
appear  in  connection  with  the  sections  that  they  have  prepared  for  me. 

t  In  England  all  analytical  chemists  are  styled  assayers.  In  the  United 
States  the  term  assayer  is  applied  to  those  chemists  who  busy  themselves 
chietlv  with  the  determination  of  the  valuable  metals. 


may  still  contain  an  appreciable  amouLit  of  copper.  Hence,  acid 
slags  should  frequently  be  treated  by  fusion  with  alkaline  carbon- 
ates, as  is  customary  in  analyzing  silicates. 

The  assayer  of  the  present  day  will  find  it  convenient  to  be 
thoroughly  familiar  with  the  three  methods  that  are  sufficiently 
accurate  and  concise  to  be  practically  employed  for  the  quantita- 
tive determination  of  copper  in  all  classes  of  material.  These 

1.  The  electrolytic  assay. 

2.  The  improved  cyanide  assay. 

3.  The  iodide  assay. 

To  which  may  be  added,  under  exceptional  conditions, 

4.  The  colorimetric  assay. 

5.  The  Lake  Superior  fire-assay. 

I  am  aware  that  the  statement  that  the  first  and  third  of  these- 
methods  are  practically  equal,  in  scope  and  exactness,  will  be  re- 
ceived with  incredulity  by  many  experienced  chemists.  It  took 
me  some  years  to  learn  that  the  improved  cyanide  method  gave 
results  almost  equal  to  the  battery  assay,  when  executed  with  equal 
skill;  and  it  is  only  on  a  recent  visit  to  England  that  I  began  to 
appreciate  the  extent  to  which  the  ioilide  assay  has  replaced  the  elec- 
trolytic method  in  that  conservative  land. 

In  some  of  the  largest  and  most  carefully  conducted  works  in 
England,  and  especially  in  one  smelter,  that  has  a  very  extensive 
electrolytic  refining  plant  of  its  own,  the  iodide  assay  is  employed 
to  the  exclusion  of  all  other  methods,  and,  as  I  was  assured  by  the 
chemist  in  charge  of  the  laboratory,  with  more  satisfactory  results 
than  the  battery  assay. 

At  least  two  among  the  best  public  laboratories  in  this  country 
are  now  making  all  their  copper  determinations  by  this  method, 
and  I  have  letters  from  the  principals  of  each  of  these  offices, 
stating  that  they  intend  using  it  in  place  of  the  battery  assay. 

The  battery  jars  are  always  a  nuisance;  and  even  where  a  con- 
stant current  can  be  obtained  from  the  electric-light  wires,  the 
iodide  assay  seems  to  be  preferred  by  several  chemists  who  have  a 
large  amount  of  work  to  do,  and  who  have  given  it  a  fair  trial. 

Hence,  I  feel  that  it  will  be  useful  to  the  profession  to  give  the 
details  of  the  operation  at  length,  both  as  practised  in  England, 
and  as  modified  by  one  of  our  most  experienced  American, 



This  is  suited  to  uearly  every  class  of  material  and  every  per- 
centage of  copper,  from  the  highest  to  the  lowest,  and,  owing  to 
its  ease  of  execution  and  extreme  accnracy,  has  already  largely 
supplanted  the  ordinary  analytic  methods,  and  bids  fair  to  do  so 
altogether  in  all  important  cases.  Among  those  assayers  who  do 
not  yet  practise  it,  there  seems  to  be  an  impression  that  it  is  diffi- 
cult of  execution,  and  in  several  cases  under  the  author's  observa- 
tion it  has  been  abandoned  after  a  few  futile  efforts.  In  these 
instances  there  must  have  been  some  direct  violation  of  the  laws 
governing  the  generation  and  transmission  of  electricity — it  being 
always  the  battery  that  was  complained  of — and  as  a  similar  though 
usually  a  much  more  extensive  and  complicated  form  of  battery  is 
under  the  charge  of  every  telegraph  operator,  the  disappointed 
assayer  should  feel  encouraged  to  persist. 

Messrs.  Torrey  &  Eaton  have  also  investigated  the  eifect  of  vari- 
ous substances  upon  the  battery  assay,  and  have  arrived  at  the  fol- 
lowing results,  which  are  not  quite  so  favorable  as  the  author's 
experience  in  practice  has  been:* 

^*  Silver,  when  present  in  any  considerable  proportion — from  1 
to  3  per  cent. — gives  too  high  a  result.  It  should  always  be  re- 
moved by  hydrochloric  acid. 

'■^Bismuth.,  even  when  present  in  small  quantity — \  per  cent. — 
is  partly  or  wholly  precipitated  with  the  copper,  and  must  conse- 
quently be  determined  analytically  in  the  deposit.  A  solution  of 
.970  gram  copper,  .030  gram  bismuth,  gave  97.9  per  cent,  copper 
instead  of  97  per  cent. 

''^  Lead,  derived  from  tiie  resolution  of  sulphate  of  lead  (if  pres- 
ent) by  the  wash-water,  is  partially  precipitated  with  the  copper. 
This  applies  only  to  large  percentages  of  lead. 

^^  Zinc  and  Nichel  do  not  interfere  in  quantities  up  to  30  per 

^^  Arsenic  precipitates  partly  tvith  the  copper,  and  not  after  it,  as 
has  been  supposed.  It  gives  a  bright  deposit,  but  may  be  found 
in  considerable  quantity  in  the  precipitate,  before  the  solution  is 
free  from  copper.  After  complete  precipitation  of  the  copper, 
therefore,  the  deposit  should  be  titrated  with  cyanide  of  potassium." 

The  following  description  has  been  written  for  me  by  Mr.  Francis 

*  Mr.  Sperry's  experience  shows  that  with  proper  precautions  these  unfavor- 
able results  may  be  completely  avoided 


L.  Sperry,  analytical  chemist,  and  for  five  years  chemist  to  the  Cai!- 
adian  Copper  (Jorapauy,  at  Siulbnry,  Ontario. 

The  scheme,  as  given,  is  intended  to  present  the  details  in  as 
practical  and  concise  a  manner  as  possible  without  going  beyond 
the  province  of  this  work.  Those  who  desire  to  study  more  closely 
the  electrolysis  of  other  metals,  and  also  the  treatment  of  copper 
in  oxalate  solutions  which  can  advantageously  be  made  use  of,  are 
referred  to  the  admirable  work  of  Dr.  Classen  on  "Quantitative 
Chemical  Analysis  by  Electrolysis,"  and  also  "Electro-Chemical 
Analysis,"  by  Prof.  Edgar  F.  Smith. 


Of  the  various  methods  the  chemist  has  in  hand  for  the  determi- 
nation of  copper,  the  electrolytic  method  presents  some  advantages 
■over  other  recognized  forms.  It  permits  of  reliable,  clean,  and 
rapid  work,  and  enables  the  chemist  to  remove  copper  from  a  solu 
tion  completely,  in  the  presence  of  other  metals,  which  may  subse- 
■quently  be  determined  in  the  same  solution. 

The  requirements  are  clean  platinum  cathodes  and  anodes  and  a 
uniform  current  of  electricity  of  known  strength. 

Take,  for  a  weighed  sample,  one-half  a  gram  copper  matte,  one 
or  two  grams  copper  ore,  depending  on  the  richness  of  the  ore,  and 
two  or  three  grams  for  slag. 

In  preparing  the  samples  they  should  be  passed  through  an  80 
mesh  sieve.     Weigh  out  on  an  accurate  chemical  balance. 

After  weighing  the  samples  in  duplicate  on  watch  glasses,  trans- 
fer carefully  to  No.  2  beakers,  slightly  moisten  with  cold  distilled 
water,  add  2r>  c.c.  strong  nitric  acid  and  10  to  15  drops  of  strong 
sulphuric  acid.  The  beakers  should  be  covered  with  watch  glasses 
and  set  on  the  sand  bath,  where  they  are  heated  until  the  nitrons 
acid  fumes  have  all  passed  ofE  and  the  sample  is  in  solution.  Wash 
the  watch  glasses  down  into  the  beaker,  and  evaporate  the  solution 
to  dryness. 

When  choking  white  fumes  appear,  set  the  beakers  one  side  to 
cool.  The  copper  is  now  in  the  form  of  sulphate.  Moisten  the 
mass  in  the  beakers  with  dilute  nitric  acid  (1.'20  sp.  gr.),  using 
about  6  or  T  c.c. ;  add  4  drops  of  sulphuric  acid,  40  c.c.  of 
water,  and  heat  on  sand  bath  until  the  mass  is  in  solution. 
Filter  ofE  the  insoluble  matter  (which  should  be  examined  to  see 
if  there  may  be  copper  left  in  tlie  residue  undissolved), reserving 



tlie  filtrate  in  a  No.  1  beaker.     The  solutiou  is  uow  ready  to  be 

The  electrical  energy  necessary  to  electrolyze  a  copper  solutiou 
is  furnished  by  various  batteries  of  reliable  manufacture.  If  there 
is  an  electric  light  plant  at  hand,  the  wire,  properly  insulated,  can 
be  run  through  the  laboratory  and  by  means  of  a  resistance  coil 
the  current  can  be  reduced  in  strength  sufficiently  to  permit  of 
quantitative  electrolytic  determinations.  The  Grenet,  Gravity,  or 
Grove  cell  batteries  will  be  found  well  adapted  for  generating  the 
necessary  strength  of  current  also.  The  Grenet  cell  loses  its  in- 
tensity after  long  use.     The  Gravity  cell  is  very  likely  to  act  un- 

FiG.  13. — Rack  for  Battery  Assay. 

satisfactorily  on  account  of  local  action  setting  in,  causing  polari- 
zation of  the  electrodes,  and  the  electrical  energy  ceasing  entirely. 
The  Grove  cell  requires  more  care  than  either  of  the  others  spoken 
of,  but  the  electromotive  force  is  certain  to  act  for  as  long  a  time 
as  is  necessary  for  the  deposition  of  the  metal,  as  the  copper  solu- 
tions are  set  on  the  battery  at  night  and  removed  on  the  following 
morning,  usually. 

It  is  best  to  have  a  surplus  of  electrical  energy  for  the  electroly- 
sis, although  too  strong  a  current  must  be  guarded  against. 

Three  Grove  cells,  freshly  made  up,  will  furnish  current  suffi- 
cient to  electrolyze  six  to  eight  copper  solutions,  none  of  wliicli- 
contain  more  than  .5  gram  copper  in  1  gram  of  sample. 



Four  cells  will  electrolyze  eight  to  ten  solutions,  and  five  cells, 
ten  to  twelve  solutions. 

A  convenient  arrangement  for  supporting  the  cathodes  and 
anodes  for  as  many  as  twelve  simultaneous  determinations  of  cop- 
per is  shown  by  the  illustration  (Fig.  13). 

The  rods  are  f  inch  square  by  39  inches  long.  Holes  for  the 
insertion  of  anode  and  cathode  rods  are  3^  inches  apart  and  j\ 
inch  in  size,  while  through  the  side  of  the  brass  rods  a  milled 
screw  sets  against  a  flexible  brass  shoe,  which  binds  the  cathode 
and  anode  platinum  rods  securely  in  position.  The  brass  rods, 
^  inch  apart,  are  supported  on  glass  pillars,  and  can  be  raised  or 
lowered  as  required. 



Fig.  14 

Fig.   15. 

Fig.  16. 

The  most  convenient  form  of  cathode  is  a  plain  platinum  cylin- 
der 2^  inches  long,  1  inch  diameter,  and  the  rod  that  supports  it 
is  4^  inches  long.     It  weighs  about  16  to  18  grams  (Fig.  14). 

These  cathodes  may  seem  large,  hut  for  general  class  of  work  on 
high  and  low  grade  copper  ores  and  mattes  they  will  be  found  a 
very  convenient  size,  as  they  offer  a  large  surface  for  the  deposition 
of  copper,  whereas,  if  they  were  smaller,  the  copper  would  fre- 
quently be  deposited  in  spongy  form,  and  there  would  be  loss  in 

The  anodes  are  platinum  wire  of  size  to  fit  y'g-inch  hole,  made 
in  the  form  of  a  concentric  circle,  from  the  center  of  which  the 
rod  stands  out  7  inches  (Fig.  15). 

The  diameter  of  the  coil  is  1  inch.  By  this  arrangement  of  the 
anode  there  is  a  uniform  evohition  of  gas  througliout  the  solution, 


aud  the  inside  as  well  as  the  outside  of  the  cathode  is  evenly  elec- 
troplated with  copper. 

The  cathode  should  not  be  completely  immersed  in  the  solution 
to  be  electrolyzed.  When  it  is  supposed  that  all  the  copper  is  de- 
posited, immerse  the  cathode  deeper  in  the  solution  and  let  the 
current  run  one-half  hour  longer.  Any  deposition  on  the  clean 
surface  will  show  at  once  that  copper  remains  still  in  the  solution. 
If  the  copper  is  all  deposited,  loosen  the  anode  and  carefully  re- 
move it  and  the  beaker.  Wash  the  cathode  quickly  into  a  clean 
No.  3  beaker  with  distilled  water,  immerse  in  pure  alcohol,  and 
gently  ignite  in  flame  until  dry.  The  copper  should  be  a  pink  rose 
color.     Weigh  as  soon  as  cooled  to  the  temperature  of  the  room. 

The  current  should  not  be  passed  through  the  solution  longer 
than  is  necessary  to  effect  the  complete  deposition  of  the  copper, 
as  secondary  reactions  are  liable  to  set  in. 

When  there  has  been  a  separation  of  copper  in  a  nitric  acid  solu- 
tion alone,  the  solution  should  be  siphoned  off  into  a  clean  beaker 
without  interrupting  the  current,  and  the  cathode  washed  with 
pure  water,  otherwise  the  nitric  acid  will  dissolve  some  of  the 
deposited  copper  into  the  solution  again.  Too  much  nitric  acid 
will  keep  the  copper  in  solution.  Too  much  sulphuric  acid  will 
cause  the  copper  to  deposit  in  spongy  form. 

Too  strong  a  current  will  cause  loss  by  too  great  evolution  of 
oxy-hydrogen  gas,  the  copper  will  deposit  dark  colored,  and  if  zinc 
is  presentjit  will  deposit  on  the  copper. 

By  using  deep  beakers  (Fig.  16),  there  will  be  scarcely  a  per- 
ceptible loss  of  solution  by  the  evolution  of  gas,  as  the  sides  of  the 
beaker  should  be  carefully  washed  down  half  an  hour  before  re- 
moving the  cathode  to  weigh. 

The  secondary  reactions  to  be  guarded  against  in  passing  the 
current  longer  tl'au  is  necessary  to  deposit  the  copper,  and  also  in 
having  the  solution  of  proper  strength  of  acid,  are  the  conversion 
of  the  nitric  acid  into  ammonia  and  the  formation  of  ammonia 
sulphate,  so  that  if  the  deposition  of  copper  were  done  in  the 
presence  of  iron  and  zinc,  these  metals  wonld  be  deposited  on  the 
cathode  as  hydrated  oxides. 

With  the  conditions  described  above  conformed  to,  copper  is 
completely  deposited  and  removed  from  solutions  containing  iron, 
alumina,  manganese,  zinc,  nickel,  cobalt,  chromium,  cadmium, 
lime,  barium,  strontium. and  magnesinra. 

In  the  laboratory  of  the  writer  it  has  been  customary  to  make 



electrolytic  separations  of  copper  and  nickel  daily  for  the  past  five 
years,  and  in  every  case  with  unvarying  accuracy.  The  copper 
was  removed  completely  in  the  presence  of  nickel,  iron  and  zinc, 
and  these  elements  subsequently  determined  in  the  same  solution. 

Examples  could   be  given  ad  infinitum,  but  a  few  will  be  given 
of  the  most  characteristic. 

f    1  gm. 

Slag -\  Copper, 

[  Nickel, 

r    1  gni. 

Ck)PPER  Ore -j  Copper, 


r    igm. 

Nickel  Ore -j  Copper, 

L  Nickel, 

Sample  taken. 

1  2 

0.42^  0.42^ 

0.41^  0.40)« 

Sample  taken. 
1  2 

8.44^  8.43JS 

3.33Jf  3.35« 

Sample  taken. 
1  2 

1.23*  1.24i« 

8.62*  8.635t 

Copper  Matte. 

f    -5  gm. 


^  Copper, 


Sample  taken. 

1  2 

33.45*        33.46* 

15.75*        15.783« 

Nickel  Matte 

r    .5gm. 

j  Copper, 

Sample  taken. 

1  2 

16.76*        16.78){ 

21.23*        21.25* 

In  each  case  the  nickel  was  determined  electrolytically  in  the 
same  solution  from  which  the  copper  was  removed. 

By  carefully  noting  what  are  the  best  conditions,  as  there  is  a 
certain  limit  within  which  variation  in  the  treatment  of  miscellane- 
ous samples  is  warranted,  most  any  novice  will  find  electrolysis  a 
simple  and  accurate  method  for  the  estimation  of  copper. 

Having  noticed  a  novel  and  most  cheap  and  convenient  appa- 
ratus for  electrolytic  assaying  in  the  laboratory  of  Messrs.  Von 
Schulz  &  Low,  in  Denver,  Mr.  Low  has  been  kind  enough  to 
furnish  me  with  the  following  description  of  the  same,  and  of  his 
method  of  using  it. 




This  apparatus  was  devised  as*  the  result  of  experiments  under- 
taken  with  a  view  of  accelerating  the  electro-deposition  of  copper 
in  analytical  work.  It  possesses  the  merit  of  simplicity  and  rapidity 
of  action,  requiring  much  less  care  than  a  battery,  and  depositing 
the  copper  in  good,  reguliue  condition  in  about  one-third  of  the 

It  consists  of  a  small  crystallization  dish,  or  beaker,  about  two 
iuclies  in  diameter,  in  which  a  stout  glass  tube,  A,  is  held  by  the 
support  B  (Fig.  17).  Two  short  glass  tubes,  passing  through 
suitably  shaped  pieces  of  cork,  are  drawn  together  with  rubber 
bands  so  as  to  hold  the  large  tube  firmly,  and  yet  permit  of  its 
being  easily  raised  or  lowered  as  required.  The  lower  end  of  the 
tube  A  is  ground  squarely  across  and  covered  with  a  parchment- 
paper  diaphragm,  which  is  attached  as  follows:  A  piece  of  stout 
parchment  paper,  about  two  inches  square,  is  thoroughly  wetted, 
and  the  superfluous  moisture  wiped  off.  If  the  paper  is  thin,  two 
thicknesses  may  be  used.  The  paper  is  now  pressed  over  the  end 
of  the  tul)e,  and  secured  tightly  in  place  with  a  rubber  band  wound 
around  as  near  the  end  of  the  tube  as  practicable.  The  loose  edges 
of  the  paper  are  cut  away  close  to  the  rubber  with  a  sharp  knife, 
and  the  joint  is  made  water-tight  by  means  of  a  little  melted  paraf- 
fine,  applied  with  a  brush.  It  requires  but  a  few  minutes  to 
attach  a  diaphragm  that  will  serve  for  several  determinations. 
The  tube  A  is  provided  at  the  top  with  a  stopper,  through  which 
passes  an  amalgamated  zinc  rod  C,  reaciiing  nearly  to  the  bottom. 
A  small  groove  is  made  in  the  side  of  the  stopper  to  permit  the 
escape  of  gas. 

In  the  bottom  of  the  outer  cup  rests  a  platinum  electrode  con- 
sisting of  a  circular  base,  about  one  and  five-eighths  inches  in 
diameter,  supporting  a  series  of  four  concentric  walls,  about  one- 
half  inch  high.  D  is  a  stout  platinum  wire  extending  up  out  of 
the  cup.  There  is  attached  to  its  lower  end  where  it  joins  the 
body  of  the  electrode,  a  piece  of  platinum  foil  reaching  up  a  short 
distance,  to  increase  the  depository  surface,  and  prevent  a  powdery 
deposit  on  the  wire.  The  entire  electrode  is  made  of  thin  plat- 
inum foil,  soldered  with  gold.  It  weighs  about  eight  grams,  and 
exposes  (including  both  sides),  about  twenty  square  inches  of  sur- 
face.    The  zinc  rod  passing  through  the  stopper  must  be  amalga- 

J.-,,;     17  — Low's   ET.F.fTTTOI.YTir   Appakatus. 


mated  with  mercury,  aud  a  short  copper  wire  is  provided  to  cou- 
uect  the  zinc  with  the  wire  D  of  the  platinum  electrode. 


The  solution  for  electrolysis  should  not  contain  more  than  one 
gram  of  copper,  which  may  be  present  either  as  sulphate  or  nitrate. 
Tlie  free  acid  should  be  neutralized  with  ammonia,  aud  then  the 
sohition  made  acid  with  a  slight  excess,  say  one  c.c,  of  strong 
nitric  acid.  The  apparatus  being  taken  apart,  the  copper  solution 
is  placed  in  the  outer  cup,  which  it  should  be  made  to  till  to  the 
depth  of  about  three-quarters  of  an  inch.  The  phitinum  electrode, 
having  been  ignited,  cooled,  and  weighed  in  the  usual  manner,  is 
now  put  in  place.  The  tube  A  is  now  about  half  filled  with  a  cold 
mixture  of  one  part  strong  sulphuric  acid  and  four  parts  water, 
anil  the  stopper  and  zinc  rod  inserted.  Finally,  the  tube  is  placed 
in  the  holder  and  adjusted  in  the  cup  so  that  the  diaphragm  just 
rests  upon  the  surface  of  the  copper  solution.  The  latter  should 
be  cold,  or  nearly  so.  Upon  connecting  the  zinc  with  the  wire 
D,  the  deposition  of  the  copper  begins  at  once,  and  is  finished  in 
from  one  to  two  hours,  according  to  the  amount  of  copper  present. 
The  apparatus  requires  but  little  watching.  It  is  well  to  raise  and 
lower  the  platinum  electrode  slightly,  to  keep  the  solution  well 

After  the  solution  has  become  perfectly  colorless  and  the  opera- 
tion appears  to  be  finished,  the  platinum  electrode  is  removed, 
dried,  and  weighed,  as  follows:  Without  disconnecting  them,  the 
tube  B  and  the  electrode  are  lifted  together  from  the  glass  cup, 
and  while  the  tube  is  returned  to  its  place,  the  electrode  is  immersed 
in  a  beaker  of  water  alongside.  If  considered  desirable,  the  elec- 
trode is  rinsed  with  a  little  water  as  it  is  taken  from  the  cup,  but 
the  copper  solution  is  usually  so  exhansted  that  the  few  drops  left 
adhering  to  the  electrode  are  of  no  importance.  The  electrical 
connection  is  now  broken  by  detaching  the  wire  W,  and  the  elec- 
trode is  further  washed,  first  with  water  aud  then  with  alcohol, 
and  finally  dried  in  the  usual  way.  It  is  now  ready  for  weighing, 
to  facilitate  which,  a  leaden  counterpoise  may  be  made,  weighing 
a  trifle  less  than  the  electrode,  so  that  the  difference  may  be  quickly 
noted  by  the  rider  of  the  balance;  and  accordingly,  in  weighing 
the  deposited  copper,  it  is  simply  necessary  to  deduct  the  amount 
previously  determined  by  the  rider.     The  electrode,  after  weighing, 


is  cleaned  with  nitric  acid,  and  again  ignited  and  weighed,  and 
re])laced  in  the  battery  for  a  short  time,  to  see  if  any  traces  of  cop- 
per remain  in  the  solution. 

Arsenic  and  antimony,  if  present  in  the  copper  solution,  axer- 
cise,  of  course,  their  usual  interference.  In  the  treatment  of  aii 
ore,  the  copper  is  best  obtained  in  the  metallic  state,  and  then  re- 
dissolved  in  a  little  nitric  acid.  It  may  be  precipitated  oij  zinc 
from  a  sulphuric  acid  solution  in  the  usual  way,  but  the  writer 
prefers  to  precipitate  it  on  a  few  strips  of  sheet  aluminum,  from  a 
boiling  solution  containing  50  c.c.  of  water  and  10  c.c.  of  strong 
sulphuric  acid.  The  copper  is  all  thrown  down  in  aboat  five  min- 
utes, and  may  be  easily  washed,  pouring  the  washings  through  a 
iilter,  and  redistolved  in  nitric  acid,  which  does  not  attack  the 
aluminum.  There  appears  to  be  no  simple  way  to  remove  anti- 
mony, but  arsenic  may  be  sufficiently  got  rid  of  as  follows: 

Evaporate  the  original  nitric  acid  solution  of  the  ore,  in  a  small 
flask,  nearly  to  dryness,  and  add  5  c.c.  of  strong  hydrochloric 
acid.  Boil  till  half  the  hydrochloric  acid  is  gone,  and  then  add 
about  2  c.c.  of  a  solution  of  2  grams  of  sulphur  in  10  c.c.  of 
bromine.  Again  boil  for  half  a  minute,  and  then  add  10  c.c.  of 
strong  sulphuric  acid,  and  heat  strongly  until  the  latter  is  boiling 
freely.  As  the  acids  and  bromine  go  off,  so  does  the  arsenic. 
What  little  may  remain  will  not  come  down  during  the  subsequent 
deposition  of  the  copper. 

The  results  obtained  by  the  above  apparatus  do  not  differ  from 
those  yielded  by  the  ordinary  battery  method,  and  the  time  required 
does  not  exceed  an  hour  and  a  half  in  any  case. 


This  well-known  and  rapid  method  depends  upon  the  power 
possessed  by  an  aqueous  solution  of  potassium  cyanide  to  decolorize 
an  amraoniacal  solution  of  a  copper  salt,  and  is,  under  proper  con- 
.'(itions,  quite  accurate  enough  for  ordinary  mill-work  on  familiar 
ores.     These  conditions  are  as  follows: 

(a)  The  use  of  measured  and  constant  amounts  of  acid  and 

(h)  The  cooling  of  the  amraoniacal  copper  solution  to  nearly 
the  temperature  of  the  surrounding  atmosphere  before  titration. 

(c)  The  intimate  mixture  of  the  cyanide  solution,  as  it  drops 
from  the  burette,  with  the  copper  solution,  and  a  sufficient,  though 
accurately  limited,  time  in  which  to  accomplish  its  bleaching  action^ 


(d)  The  establishment  of  a  certain  fixed  shade  of  pink  at  the 
standardizing  of  the  cyanide  solution,  to  which  all  subsequent 
assays  mitst  be  as  closely  as  possible  approximated  in  color  for  the 
finishing  point.  This  renders  it  impossible  for  any  chemist  to 
work  with  another  person's  solution  until  he  has  first  standardized 
it  himself,  and  determined  its  strength  according  to  his  own 

The  absence  of  any  considerable  amount  of  lime,  zinc,  arsenic 
and  antimony,  whose  presence  has  long  been  known  to  seriously 
vitiate  results,  though  exactly  to  what  extent  was  not  demon- 
strated, until  a  series  of  experiments  on  this  point  was  carried  out 
in  1883  under  the  direction  of  the  author,  and  still  more  recently 
by  Torrey  &  Eaton. 

From  a  long  list  of  results,  some  of  them  even  contradictory, 
the  following  deductions  were  drawn: 

The  presence  of  zinc  in  quantities  below  4^  per  cent,  has  no 
perceptible  influence  on  results. 

Five  per  cent,  of  zinc,  in  a  siliceous  ore  of  copper,  containing 
no  other  metals  except  iron,  caused  a  constant  error  on  the  plus 
side  of  about  0.32  per  cent.,  which  increased  in  a  tolerably  regular 
ratio  with  an  increased  percentage  of  zinc. 

After  eliminating  a  few  results  that  varied  very  greatly  and  un- 
accountably from  all  others,  an  average  of  about  six  determinations 
of  each  sample  yielded  the  following  figures.  The  ore  just  de- 
scribed was  used  in  every  case,  and  the  zinc  added  in  the  shape  of 
a  carefully  determined  sulphide,  allowances  being  also  made  for 
the  increase  in  the  weight  of  the  ore  sample  resulting  from  this 
addition  of  foreign  matter. 

Ore  free  from  zinc,  11.16  per  cent,  copper. 

No.  1  with  4    per  cent,  zinc,  11.46 

"  11.55 

"  11.72 

"  12.1 

"  13.2 

"  13.3 

"  13.9 

"  13.8 

The  presence  of  arsenic  and  antimony  in  much  smaller  propor- 
tions— 1  per  cent,  or  less — may  cause  errors  on  both  plus  ard 
minus  sides  to  the  extent  of  one-half  a  per  cent,  or  more,  and  in 
larger  quantities  will  generally  render  the  test  totally  unreliable. 

2     ' 

'     H 

3    ' 

'      5 

4     ' 

•     6 

5     ' 

'     8 

6     ' 

'    10 


'    15 

8     • 

'    20 


Another  indispensable  and  oft-neglected  precaution  is  the  testing 
of  the  precipitate  of  hydrated  oxide  of  iron  caused  by  the  addition 
of  ammonia  to  the  original  solution.  This  bulky  precipitate,  es- 
pecially in  the  case  of  mattes  and  highly  ferruginous  ores,  will 
retain  a  considerable  amount  of  copper  which  even  the  most  care- 
ful wasliing  will  not  remove,  but  which  may  be  speedily  deter- 
mined by  dissolving  the  precipitate  in  the  smallest  possible  quan- 
tity of  muriatic  acid,  saturating  with  ammouia,  and  again  titrating 
if  any  blue  coloration  is  produced.  The  following  results,  taken 
from  the  notebook  of  a  busy  chemist,  who  had  never  been  aware 
of  this  possible  source  of  error  until  accidentally  mentioned  to 
him  by  an  assayer  in  the  employ  of  the  writer,  and  who  at  once 
instituted  careful  experiments  to  ascertain  the  probable  extent  of 
the  mistakes  that  he  had  made  while  acting  as  assayer  to  large 
smelting  works,  give  some  idea  of  the  serious  discrepancies  that 
may  arise  from  the  non-observance  of  this  precaution: 

Without  resolution  of  "With  resolution  of 

Character  of  sample.  precipitate  precipitate 

No.  1,  pyritous  ore 21.2  per  cent,  copper  23.7  per  cent,  copper 

•'  2,bornite... 37.8  "  "  42.4 

"  3,  cupola  luatte 27.7  '•  "  31.2         "  " 

"  4,  reverbatory  matte 46.4  "  "  47.4         "  " 

"  5,  blue  metal 57.7  "  "  58.2         "  " 

"  6,  white    metal 74.7  "  "  75.2         "  " 

"  7.  regule 86.2  "  "  86.5 

"  8,  blister  copper 97.3  "  "  97.2 

As  might  be  expected,  the  greatest  discrepancies  exist  in  con- 
nection with  those  samples  containing  the  largest  amounts  of  iron, 
and  decrease  to  nothing  as  the  iron  contents  diminish. 

In  the  absence  of  the  injurious  elements — zinc,  arsenic,  anti- 
monv — the  cyanide  assay  is  sufficiently  accurate,  and,  from  its 
simplicity  and  rapidity  of  execution,  it  is  peculiarly  adapted  to 
the  daily  working  assays  from  the  mine,  smelter  and  concentrator. 
In  fact,  it  is  the  mainstay  of  the  overcrowded  metallurgical 
assayer,  and  can  be  used  for  nearly  every  purpose,  except  for  the 
bnving  and  selling  of  ores  and  copper  products,  and  for  the  deter- 
mination of  very  minute  quantities  of  copper  in  slags.  It  is  fre- 
quently employed  with  satisfaction  for  the  last-named  purpose,  a 
much  larger  amount  than  usual  being  taken,  in  order  to  obtain 
a  solution  sufficiently  rich  in  copper  to  exhibit  a  reasonable  degree 
III  color.     Messrs.  Torrey  &  Eaton  have  published  additional  in- 

THE    SAilPLIXG    AXD    AiSSAYIXG    OF    COPPER.  00 

vestigations  of  great  value  on  the  effect  of  various  snbstauces  upon 
the  accuracy  of  the  cyanide  niethoJ.  (See  Engineering  and  Mining 
Journal,  May  9,  I880.) 

In  their  experiments  they  employed  a  cyanide  solution  capable 
of  showing  one-thirtieth  of  one  per  cent,  of  copper,  and  took  every 
precaution  to  have  all  conditions  identical  during  the  various  tests; 
all  solutions  titrated  being  of  the  same  degree  of  strength. 

Silver  and  Bismuth. — A  solution  was  made  of  the  following 

Copper 550  gram. 

Bismutli 200      " 

Silver    350      " 

The  silver  was  precipitated  with  hydrochloric  acid,  and  ammonia 
added  after  filtering  and  washing.     Two  titrations  gave: 

No.  1 54.90  per  cent,  copper 

No.  2 54.85         •'  "         instead  of  55  per  cent. 

These  results  show  that  a  solution  containing  the  very  unusual 
proportion  of  20  per  cent,  of  bismuth  and  25  per  cent,  of  silver 
can  be  titrated  to  within  0.1  per  cent,  of  its  value  in  copper. 

Lead. — This  metal,  being  a  common  element  in  copper  ores  and 
alloys,  was  introduced  into  a  copper  solution  in  the  following  pro- 

Copper    • 200  gram. 

Lead 800 

After  adding  ammonia  and  allowing  the  lead  precipitate  to  sep- 
arate for  two  or  three  hours,  it  was  titrated,  giving  20.28  per  cent, 
of  copper,  instead  of  20  per  cent.  Messrs.  Torrey  &  Eaton,  there- 
fore, believe  that  the  amount  of  lead  commonly  present  in  ores — 
from  5  to  40  per  cent. — would  not  injuriously  affect  the  ojjeration. 

Arsenic. — Torrey  &  Eaton  titrated,  without  filtering,  a  solution 
containing  .600  gram  arsenic,  .400  gram  copper,  finding  39.8  per 
cent,  instead  of  40  per  cent.  Therefore  any  ordinary  amount  of 
arsenic — from  5  to  15  per  cent. — would  seem  to  have  no  injurious 

Ammonia  and  hydrochloric  acid,  when  indiscriminately  used, 
were  found  by  Messrs.  Torrey  &  Eaton  to  cause  serious  errors,  the 
results  being  influenced  to  the  extent  of  from  |  to  1  per  cent,  by 
any  large  excess  of  either. 


Lime  in  large  quantity  was  found  to  confuse  results. 
Magnesia  had  no  efifect  whatever. 

low's   modified   cyanide   assay.* 

The  cyanide  method,  preceded  by  the  separation  of  the  copper 
from  interfering  substances,  and  more  or  less  modified  by  different 
operators,  is  the  one  in  common  use  in  Colorado  for  technical 
work.  The  writer  has  adapted  a  modification  of  his  own,  in  which 
the  preliminary  precipitation  of  the  copper,in  the  metallic  state,  is 
effected  by  aluminum  instead  of  the  customary  zinc.  The  object 
is  to  obtain  a  copper  that  is  unquestionably  free  from  zinc,  as  it  is 
found  that  sometimes,  when  an  excess  of  zinc  is  employed,  some 
of  that  metal  is  retained  by  the  copper,  causing  too  high  a  result. 
Of  course,  such  an  error  is  accidental  and  unnecessary,  but  it  is, 
nevertheless,  sufficiently  frequent  to  indicate  the  value  of  a  scheme 
in  which  it  cannot  possibly  occur.  The  copper  is  eventually  ob- 
tained in  a  blue,  ammouiacal  solution,  and  its  amount  is  estimated 
from  the  quantity  of  standard  solution  of  cyanide  of  potassium 
required  to  discharge  the  blue  color,  as  in  the  ordinary  cyanide 
assay.  The  results  of  the  cyanide  titration  are  exact  if  certain 
conditions  are  always  maintained.  It  is  found  that  for  the  same 
amount  of  copper: 

1.  A  concentrated  solution  requires  more  cyanide  for  decolora- 
tion than  a  dilute  solution. 

'l.   A  hot  solution  requires  less  cyanide  than  a  cold  one. 

0.  In  any  case  when,  from  a  rapid  addition  of  cyanide,  the  color 
has  become  rather  faint,  it  may,  by  simple  standing,  continue  to 
fade,  and  perhaps  entirely  disappear. 

•4.  If  the  amount  of  cyanide  added  is  insufficient  to  effect  com- 
plete discharge  of  color,  even  after  allowing  the  copper  solution  to 
stand  for  several  minutes,  the  titration  may  then  be  finished  with- 
out alteration  of  the  final  result. 

From  the  foregoing  facts  it  is  evidently  necessary,  in  order  to 
obtain  correct  results,  that  the  titrations  for  unknown  amounts  of 
copper  should  be  made  under  conditions  that  do  not  differ  materi- 
ally in  the  following  particulars  from  those  governing  the  stand- 
ardization of  the  cyanide  solution: 

1.  Temperature. 

*Mr.  A.  H.  Low  has  been  kind  enouo^b  to  send  me  tbe  following  description 
of  bis  very  convenient  and  useful  modification  of  tbe  Cyanide  Assay. 


2.  Rapidity  of  the  final  additions  of  cyanide. 

3.  Final  bulk  of  solution. 

Besides  the  physical  conditions  just  enumerated,  there  are 
chemical  conditions  that  effect  the  result,  such  as  presence  of  a 
large  amount  of  chlorides,  a  large  excess  of  ammonia,  etc.  Such 
abnormal  conditions  require  no  special  consideration,  since  they 
are  all  easily  avoided  by  following  the  niethod  to  be  described. 


Dissolve  pure  cyanide  of  potassium  in  distilled  water,  in  the 
proportion  of  21  grams  to  the  liter.  The  commercial  cyanide, 
dissolved  in  common  water,  may  be  used,  but  is  not  recommended. 
An  uncertain  quantity,  perhaps  60  grams  to  the  liter,  is  required, 
and  a  slimy  precipitate  or  residue  is  always  left,  that  must  either  be 
filtered  off,  or  allowed  to  settle,  so  as  to  decant  the  pure  liquid. 

Weigh  accurately  about  U.200  grams  of  pure  copper  foil,  and 
place  it  in  a  pear-shaped  flask  of  about  250  c.c.  capacity.  Add  5 
c.c.  of  strong,  pure  nitric  acid,  which  will  quickly  dissolve  the 
copper.  Without  boiling  off  the  red  fumes,  add  about  80  c.c.  of 
<iistilled  water  and  10  c.c.  of  strong  ammonia  water  (26  degrees 
Beaume).  Cool  to  the  ordinary  temperature,  by  placing  under  a 
tap,  or  in  cold  water.  Titrate  with  the  cyanide  solution,  in  a 
slow,  cautions  manner,  and,  as  the  end-point  is  approached,  as 
shown  by  the  partial  fading  of  the  blue  color,  add  distilled  water 
so  as  to  bring  the  solution  to  a  bulk  of  approximately  150  c.c. 
Finish  the  titration  by  careful  and  regular  additions  of  cyanide, 
finally  decreasing  to  a  drop  at  a  time,  until  the  blue  tint  can  no 
longer  be  detected  by  holding  the  flask  against  a  light-colored 
background.  It  is,  of  course,  very  essential  that  there  should  be 
no  haste  and  no  prolonged  delay  in  these  final  additions  of  cyanide. 
Simply  adopt  a  regular,  natural  manner,  that  can  easily  be  re- 
peated on  all  subsequent  titrations.  Keep  the  cyanide  solution  in 
a  glass-stoppered  bottle  (the  common  2^  liter  acid-bottle  is  conven- 
ient), in  a  cool  place  not  exposed  to  direct  sunlight.  Under  these 
circumstances  it  holds  its  strength  fairly  well,  but  still  it  gets 
weaker  from  the  decomposition  of  the  cyanide,  and  should  be 
restandardized  weekly. 


Treat  1  gram,  or  0.5  gram  if  the  material  seems  to  contain  40 
per  cent,  copper  or  over,  in  a  flask  of  about  250  c.c.  capacity,  by 


boiling  first  with  10  to  15  c.c.  of  strong  nitric  acid  to  effect  de- 
composition, and  then  with  10  c.c.  of  strong  sulphuric  acid,  to 
expel  the  nitric  acid.  In  most  cases,  ores  are  easily  decomposed, 
and  do  not  require  extremely  fine  pulverization;  the  assay  that  has 
passed  an  80-mesh  sieve  being  fine  enough.  In  the  case  of  mattes 
however,  it  is  best  to  give  an  additional  grinding,  in  an  agate 
mortar,  of  a  small  portion,  from  which  the  sample  for  analysis  is 
to  be  weighed. 

Boil  the  contents  of  the  flask  gently  during  the  decomposition 
with  nitric  acid,  and  then,  after  the  addition  of  the  sulphuric 
acid,  place  the  flask  over  a  small,  naked  flame,  and  heat  until  all 
the  nitric  acid  is  expelled,  and  the  residuary  sulphuric  acid  is 
boiling  freely  and  evolving  copious  fumes.  Remove  from  the 
flame  and  allow  to  cool.  Ores  that  are  not  decomposed  by  this 
treatment  must  be  attacked  in  some  special  manner  for  which  no 
general  directions  can  be  given.  Sometimes  the  addition  of 
hvdrochloric  acid  is  all  that  is  necessary.  It  is  advisable  in  any 
case  not  to  add  the  sulphuric  acid  until  the  ore  appears  to  be  well 

To  the  residue  in  the  flask  add  50  c.c.  of  water,  and  three  or 
more  pieces  of  sheet  aluminum,  eacli  perhaps  1^  by  ^  by  ^\  inches 
in  size,  and  heat  the  mixture  tc  boiling.  Boil  strongly  for  at 
least  five  minutes,  when  the  copper  is  ordinarily  all  precipitated. 
Eeniove  from  the  lamp,  add  25  c.c.  of  cold  water,  allow  to  settle 
a  moment,  and  then  decant  through  a  three-inch  filter,  retaining 
in  the  flask  the  aluminum  and  as  much  of  the  copper  as  possible. 
Wash  the  precipitated  copper  twice  by  decantation,  using  about 
25  c.c.  of  water  each  time,  and  pouring  through  the  filter.  Drain 
the  flask  as  completely  as  possible  the  last  time,  and  then  give  the 
filter  one  or  two  extra  rinsings.  Pour  into  the  flask  5  c.c.  of 
strong  nitric  acid,  and  shake  it  about  gently  until  the  copper  is  all 
dissolved,  the  aluminum  not  being  attacked.  Now  add  5  c.c.  of 
water  and  one  drop  of  strong  hydrochloric  acid,  and,  after  shaking 
around  for  a  moment  to  coagulate  any  chloride  of  silver,  pour  the 
mixture  upon  the  filter,  thus  dissolving  wliatever  copper  is  there, 
receiving  the  filtrate  in  a  small  beaker.  Rinse  out  the  flask,  pour- 
ing the  rinsings  through  the  filter,  and  remove  the  aluminum, 
which  is  but  little  attacked,  for  further  use.  Finally,  wash  the 
filter  thoroughly,  but  with  as  little  water  as  possible,  so  as  not  to 
obtain  too  bulky  a  filtrate,  and  then  transfer  the  solution  back  to 
the  flask  again.     Now  add   10  c.c.  of  strong  (26  degrees  Beaumo) 


ammonia  water,  aud  cool  the  solution  to  the  ordinary  temperature. 
Dilute  to  about  75  c.c,  and  titrate  with  the  standard  cyanide  cau- 
tiously until  the  blue  color  is  discharged  to  a  considerable  extent, 
and  it  is  evident  that  the  end  point  is  not  far  otf. 

The  liquid  is  now  frequently  more  or  less  cloudy  from  the  pres- 
ence of  hydrate  of  lead,  and  possibly  small  amounts  of  the  hydrates 
of  iron,  aluminum,  etc.,  and  for  accurate  work  should  be  filtered. 
If  the  titration  has  been  carried  too  far  before  filtration,  the  faint 
blue  tinge  is  liable  to  fade  completely  away,  thus  spoiling  the 
assay.  On  the  other  hand,  it  is  not  advisable  to  filter  the  amnio- 
niacal  solution  before  the  addition  of  cyanide,  as  such  a  filtered 
solution  will  frequently  develop  a  second  milkiness  during  the 
titration,  and  have  to  be  filtered  again.  Filter  the  partly  titrated 
solution  through  a  5-inch  filter.  One  washing  will  usually  suffice. 
Finish  the  titration  very  carefully  on  the  clear,  pale-blue  solution, 
precisely  as  in  the  standardization  previously  decribed.  Toward 
the  end,  dilute  with  water,  if  necessary,  so  as  to  obtain  a  final 
bulk  of  about  150  c.c. 

The  number  of  c.c.  of  cyanide  solution  required,  multiplied  by 
the  copper-value  of  one  c.c.,gives  the  weight  of  copper  contained  in 
the  amount  of  ore  taken,  from  which  the  percentage  is  readily 

If  the  amount  of  silver  present  in  an  ore  is  known,  it  need  not 
be  removed,  but  may  be  allowed  for,  on  the  basis  that  2Ag  =  Cu. 
One  per  cent,  of  silver,  or  292  ounces  per  ton  of  2,000  pounds, 
would  thus  approximately  equal  0.293  per  cent,  of  copper;  and 
100  ounces  of  silver  per  ton  would  equal  0.10  per  cent,  copper. 
Accordingly,  0.10  per  cent,  of  copper  is  to  be  deducted  for  every 
100  ounces  of  silver  ^er  ton.  Thus,  if  the  result  of  the  titration 
indicates  24.63  per  cent.  Cu,  and  the  ore  assays  250  ounces  per 
ton  in  silver,  the  true  result  for  copper  is  24.63 — 0.25,  or  24.38 
per  cent. 

None  of  the  ordinary  constituents  of  ores  interfere  with  the 
method  as  described.  Duplicates  should  easily  agree  within  0.10 
or  0.2  per  cent.,  which  answers  for  ordinary  uncommercial  tests. 


The  following  description  of  the  iodide  assay,  as  practiced  very 
largely  at  English  metallurgical  works  in  place  of  the  electrolytic 
method,  has  kindly  been  written  for  me  by  Mr.  I.  H.  Glutton, 
nssayer  and  metallurgical  chemist  to  The  Elliott's  Metal  Company, 


Limited;  Selly  Oak  Works,  Birmingham,  England;  and  Pembrey 
Copper  and  Silver  Works,  Burry  Port,  South  Wales,  with  the  per- 
mission of  Mr.  Gerard  B.  Elkington. 


This  assay,  which  may  now  be  fairly  described  as  the  standard 
English  method,  is,  in  practised  hands,  both  accurate  and  expedi- 

It  depends  upon  the  reaction  that  occurs  when  an  excess  of 
potassium  iodide  is  added  to  a  solution  of  a  cupric  salt  in  a  slightly 
acid,  or  acetic  acid,  solution,  cuprous  iodide  being  formed  and  an 
equivalent  of  iodine  set  free.  This  free  iodine  is  dissolved  in  the 
excess  of  potassium  iodide,  the  liberated  iodine  being  in  exact  pro- 
portion   to    the    copper    present.     Thus:  2CuSOi4-4KI  =  Cu2l2+ 

The  free  iodine  {i.e.,  indirectly  the  copper)  is  determined  in 
the  usual  way,  by  titrating  with  sodium  hyposulphite  (thiosul- 
phate),  using  starch  as  an  indicator.  Sodium  iodide  and  tetrathion- 
ate  are  formed. 

Thus  2Na2S203+L=-2NaI+Na2SA. 

The  exact  method  of  conducting  the  assay  is  as  follows: 

From  one-half  to  two  grams  (100  to  400  milligrams  of  copper) 
of  the  ore  or  matte,  according  to  richness,  is  weighed  into  an  8- 
ounce  flask  and  dissolved,  best  by  thorough  decomposition  of  the 
sulphides  with  nitric  acid,  and  after  taking  nearly  to  dryness,  by 
partial  evaporation  with  sulphuric  acid;  this  renders  any  lead 
insoluble  by  converting  it  into  sulphate.  The  lead  sulphate  and 
insoluble  residue  are  then  filtered  oflf,*  the  solution  being  passed 
into  a  IG-ounce  flask,  in  which  the  copper  is  precipitated  as  sul- 
phide either  by  hyposulphite  of  soda,  or  sulphureted  hydrogen. 
The  former  method  has  the  advantage  of  not  contaminating  the 
atmosphere  of  the  laboratory,  while,  in  making  a  large  number  of 
assays,  the  latter  is  more  economical,  and  even  perhaps  more  expe- 
ditious.    In  either  case,  the  sulpbides  are  washed  free  from  iron. 

If  much  iron  is  present,  the  solution,  during  titrating,  will  have 
a  reddish  color,  due  to  ferric  acetate.  This  tends  to  mask  the 
reaction,  but  reasonably  small  quantities  of  iron  do  not  interfere. 

The  sulphides  are  washed  back  into  the  flask,  and  dissolved  with 
10  c.c.  of  nitric  acid.     A  little  chlorate  of  potassium  may  also  be 

*  This  filtering  may  also  be  omitted. 


used  at  the  end,  to  assist  in  the  liberation  of  the  sulphur.  The 
assay  is  evaporated  on  the  sand  bath  to  as  near  dryness  as  possible; 
nitrous  fumes  are  blown  out,  and  the  copper  salts  dissolved  in  a 
little  hot  water.  The  solution  is  filtered  through  a  small  funnel 
over  Swedish  filter -paper,  into  a  No.  7  beaker,  sulphur  and  most 
of  the  antimony  (as  oxide)  remaining  behind,  together  with  lead 
sulphate  and  insoluble  residue,  if  the  first  filtering  was  omitted. 

Instead  of  dissolving  the  sulphides  direct,  some  chemists  dry, 
calcine,  and  dissolve  the  remaining  oxides  in  a  little  nitric  acid. 

The  solution,  which  should  not  now  much  exceed  50  c.c.  in 
bulk,  is  neutralized  with  sodium  carbonate,  and  a  slight  excess  of 
acetic  acid  added. 

Potassium  iodide  crystals  are  then  added,  the  quantity  being 
immaterial  so  long  as  there  is  enough;  about  five  or  six  grams  will 
be  the  proper  amount  in  ordinary  cases. 

The  solution,  in  which  the  copjier  now  exists  as  cuprous  iodide, 
is  of  a  yellow-brown  color,  and  sodium  hyposulphite  is  run  into  it 
from  a  burette,  the  assay  being  constantly  shaken.  The  yellow 
color  rapidly  lightens,  and,  on  the  addition  of  5  or  6  c.c.  of 
starch  solution,  strikes  a  deep  blue  color,  which  disappears  on  the 
addition  of  more  hyposulphite,  leaving  a  creamy,  white  tint,  the 
end  reaction  being  sharp  and  distinct.  The  assay  is  titrated  in 
the  cold,  to  prevent  any  possible  loss  of  iodine. 

The  hyposulphite  solution  alters  slightly  by  keeping,  and  requires 
occasional  standardizing.  Its  strength  is  such  that  500  milligrams 
of  copper  require  from  50  to  60  c.c.  (about  40  grams  of  hypo  to 
one  liter  of  water). 

The  standards  are  prepared  by  dissolving  4  grams  of  pure  elec- 
trolytic copper  in  the  smallest  possible  excess  of  nitric  acid,  dilut- 
ing to  one  liter,  and  drawing  off  50  c.c.  at  one  time,  neutralizing 
and  treating  exactly  as  the  assay. 

The  starch  solution  is  made  by  pouring  a  little  boiling  water  on 
about  one  gram  of  starch  in  a  beaker,  rubbing  the  same  to  a  thin 
paste  with  a  glass  rod,  neutralizing,  and  treating  exactly  as  the 

Personally,  I  find  that  in  neutralizing,  it  is  best  to  have  only 
the  slightest  excess  of  carbonate  of  sodium. 

Arsenic  does  not  interfere  with  the  action  or  affect  the  result. 
Bismuth  imparts  a  yellow  color  to  the  solution,  and  the  assay  is 
liable  to  be  overdrawn  before  the  addition  of  starch;  otherwise  it 
does  not  affect  the  reaction  or  the  result. 


Mr.  A.  H.  Low,  assayer  and  analytical  chemist,  of  Denver,  Col- 
orado, has  been  kind  enough  to  furnish  me  with  the  following 
description  of  certain  modifications  that  he  has  made  in  the  iodide 
assay,  that  tend  to  render  it  more  convenient  and  exact. 

Mr.  Low  says:  The  iodide  method  for  the  copper  assay,  when 
carried  out  according  to  the  followijig  modification,  devised  by 
the  writer,  appears  to  fully  equal  the  electrolytic  method  in  accu- 
racy, and  as  it  requires  very  much  less  time,  scarcely  more  than 
the  ordinary  cyanide  method,  it  is  greatly  to  be  preferred  for  the 
usual  run  of  impure  ores  and  furnace  products.  The  following 
figures  are  given  to  show  the  accuracy  of  the  hyposulphite  titration 
as  described :  On  the  basis  of  one  gram  as  100  per  cent,  there 

Taken    0.78  per  cent.  Cu found    0.79  per  cent.  Cu. 

10.30  "  "     10.30 

15.37  "  "     15.38 

"       20.31  "  "     20.30 

44.45  "  "     44.44 

46.92  "  "     46.89 

56.82  "  "     56.85 

Can  the  electrolytic  method  improve  upon  this?  No  special 
pains  were  taken  with  these  tests,  and  they  were  made  as  rapidly  as 
the  daily  technical  work.  The  scheme  devised  removes  all  ordinary 
interfering  impurities  or  renders  them  inert.  Zinc  has  not  been 
found  as  good  a  precipitant  for  the  copper  as  aluminum  for  several 
reasons,  one  of  the  principal  being  that  the  precipitated  copper  is 
frequently  contaminated  with  considerable  iron,  even  when  thrown 
down  from  strongly  acid  solutions,  and  this  iron  may  occasion 
much  subsequent  annoyance.  When  aluminum  is  used,  the  pre- 
cipitation may  be  effected  without  boiling  by  simply  adding  a  few 
drops  of  hydrochloric  acid  to  the  solution,  but  this  has  not  been 
found  so  desirable  as  the  method  described.  For  the  success  of 
the  hyposulphite  titration  it  is  absolutely  essential  that  there  be 
no  nitrate  of  copper  or  free  nitric  acid  present.  When  a  solution 
of  nitrate  of  copper  is  neutralized  in  the  cold  with  ammonia,  which 
may  even  be  added  in  large  e.^cess,  and  the  solution  is  then  re- 
acidified  with  acetic  acid,  the  mixture  behaves  toward  iodide  of 
potassium  as  though  there  were  some  nitrate  of  copper  or  free 
nitric  acid  also  present.  If,  however,  the  ammoniacal  liquid  be 
boiled  for  a  moment,  the  neutralization  appears  to  be  complete,  the 
nitric  acid  all  combining  with  the  ammonia  and  occasioning  no 
subsequent  trouble  on  the  addition  of  acetic  acid. 



Prepare  a  aolutiou  of  hyposulphite  of  sodium  containing  about 
38  grams  of  the  pure  crystals  to  the  liter,  standardize  as  follows: 
Weigh  accurately  about  0.200  grams  of  pure  copper  foil  and  place 
in  a  flask  of  about  250  c.c.  capacity.  Add  about  4  c.c.  of  strong 
nitric  acid  to  dissolve  the  copper  and  then  evaporate  down  to  1  or 
2  c.c,  avoiding  overheating  which  might  easily  convert  some  of 
the  copper  into  a  basic  salt  or  oxide.  The  operation  may  be  has- 
tened by  manipulating  the  flask  in  a  holder  over  a  small  naked 
flame.  Now  add  5  c.c.  of  water  to  dissolve  the  nitrate  of  copper, 
and  then  5  c.c.  of  strong  ammonia  water.  See  that  the  copper 
has  all  dissolved  and  the  solution  is  strongly  alkaline.  Heat  to 
boiliug  and  boil  for  about  a  minute.  This  is  absolutely  necessary 
to  insure  the  perfect  neutralization  of  the  nitric  acid  in  the  nitrate 
of  copper.  Remove  from  the  heat  and  add  6  c.c.  of  glacial  acetic 
acid,  and  then  40  c.c.  of  cold  water.  Again  see  that  all  copper 
salts  are  dissolved,  and  then  add  to  the  cool  solution  about  3  grams 
of  iodide  of  potassium  and  shake  it  about  gently  until  dissolved. 
Cuprous  iodide  will  be  precipitated  and  iodide  liberated  according 
to  the  following  reaction:  2[Cu.  3C2H302]+4KI  =  Cu2l2+4[K. 
C2H302J+2I.  The  free  iodine  colors  the  mixture  brown.  Ti- 
trate at  once  with  the  hyposulphite  solution  until  the  brown  tinge 
has  become  weak,  and  then  add  sufficient  starch  liquor  to  produce 
a  marked  blue  coloration.  Now  continue  the  titration  cautiously 
until  the  blue  tinge  vanishes.  Stop  at  the  first  decided  change, 
and  the  color  will  usually  entirely  disappear  after  standing  a  mo- 
ment. One  c.c.  of  the  hyposulphite  solution  will  be  found  to 
correspond  to  about  0.01  gram  of  copper.  The  reaction  between 
the  hyposulphite  and  iodine  is:  2[Na2S203]+2I  =  3Nal4-Na2S406. 
Sodium  iodide  and  tetrathionate  are  formea.  The  starch  liquor 
may  be  made  by  boiling  about  half  a  gram  of  starch  with  a  little 
water  and  diluting  to  about  250  c.c.  It  should  be  used  cold,  and 
must  be  prepared  frequently  for  regular  work,  as  it  does  not  keep 
very  well.  The  hyposulphite  solution  made  of  the  pure  crystals 
and  distilled  water  appears  to  be  very  stable.  The  writer  has 
never  detected  any  appreciable  variation  in  strength  during  the 
time  required  to  use  up  a  lot,  say  a  month  or  more. 


Treat  1  gram  of  tlie  ore  in  a  flask  of  350  c.c.  capacity  with  10 
c.c.  of  strong  nitric  acid   by    boiliug  nearly  or  quite  to  dryness. 


Now  add  lu  c.c.  of  strong  hydrochloric  acid  aud  agaiu  boil. 
After  boiliug  for  two  or  three  minutes,  add  10  c.c.  of  strong  sul- 
phuric acid  and  heat  strongly,  best  over  a  small  naked  flame,  until 
the  more  volatile  acids  are  expelled,  and  the  fumes  of  sulphuric 
acid  are  coming  oflE  freely.  Allow  to  cool,  aud  then  add  about  40 
c.c.  of  water  aud  heat  to  boiliug.  Filter  through  a  .3-inch  filter, 
more  particularly  to  remove  any  sulphate  of  lead,  aud  collect  the 
filtrate  in  a  beaker  about  three  inches  in  diameter.  Wash  flask  and 
residue  with  hot  water,  aud  endeavor  to  keep  the  volume  of  the 
filtrate  down  to  75  c.c.  or  less.  Place  in  the  bottom  of  the  beaker 
a  piece  of  sheet  aluminum  prepared  as  follows:  Cut  from  stout 
sheet  aluminum  a  strip  about  3  inches  long  and  1|  inches  wide, 
aud  bend  up  each  end  at  right  angles  for  about  five-eighths  of  an 
inch,  or  so  that  the  body  of  the  stri])  will  lie  flat  in  the  bottom  of 
the  beaker.  This  aluminum  may  be  used  repeatedly  as  it  is  but 
little  attacked  each  time.  Cover  the  beaker  aud  heat  to  boiling. 
Boil  strongly  for  about  six  or  seven  minutes,  when  the  copper  will 
be  all  precipitated  if  the  bulk  of  the  solution  does  not  exceed  T5 
c.c.  More  dilute  solutions  should  be  boiled  correspondingly 
longer,  as  the  sulphuric  acid  does  uot  begin  to  attack  the  alumi- 
num strongly  until  of  about  the  degree  of  concentration  recom- 
mended. Now  pour  the  liquid  in  the  beaker  back  into  the  original 
flask,  and, with  the  wash-bottle  of  hot  water,  rinse  in  also  as  much 
of  the  copper  as  possible,  leaving  the  aluminum  behind.  The 
beaker  and  aluminum,  which  may  still  retain  some  adhering  cop- 
per, are  now  temporarily  set  aside.  The  copper  in  the  flask  is 
allowed  to  settle  and  the  clear  liquid  decanted  through  a  small 
filter.  Wash  the  cop])er  two  or  three  times  by  decantation  with  a 
little  hot  water,  pouring  the  washings  through  the  filter  but  re- 
taining the  copper  as  completely  as  possible  in  the  flask.  Now 
place  the  beaker  containing  the  aluminum  under  the  funnel  and 
pour  .3  or  4  c.c.  of  strong  nitric  acid,  drop  by  drop,  over  the  filter. 
This  dissolves  whatever  copper  may  be  there  and  washes  it  into 
the  beaker.  Wash  with  a  little  hot  water  if  necessary,  but  endeavor 
to  keep  the  total  volume  of  liquid  as  small  as  possible.  Finally 
rinse  the  solution  in  the  beaker  into  the  flask  and  set  the  latter 
over  the  lamp.  As  soon  as  the  copper  has  dissolved, add  about  half 
a  gram  of  chlorate. of  potassium  to  oxidize  any  arsenic  present  to 
arsenic  acid.  This  is  very  important.  Continue  the  boiliug  until 
only  1  or  "2  c.c.  of  liquid  remain,  but  not  so  far  as  to  form  a  basic 
salt  or  the  oxide  of  copper.     Proceed  with  the  residue  precisely  as 


with  the  residue  of  nitrate  of  copper  in  the  standardization  of  the 
hvijosnlphite,  finally  calculating  the  percentage  of  copper  present 
from  the  amount  of  standard  hyposulphite  required. 

An  excess  of  iodide  of  potassium  is  not  necessary.  One  gram  of 
pure  copper  requires  5.24  grams  of  KI,  consequently  3  grams  of 
KI  are  quite  sufficient  for  anything  under  50  per  cent.  Cu,  when 
1  gram  of  ore  is  taken  for  assay.  If  the  percentage  of  copper  is 
likely  to  run  ahove  that  point,  5  grams  of  the  iodide  had  better  be 
used.  Lead  and  bismuth  form  colored  iodides,  and  if  present  in 
any  considerable  amount,  mask  the  end-point  before  adding  starch. 
They  are  otherwise  without  effect,  as  is  also  arsenic  when  oxidized 
as  described.  The  return  of  the  blue  tinge  in  the  liquid  by  long 
standing  after  titration  is  of  no  significance. 


This  is  reserved  almost  exclusively  for  the  determination  of 
minute  quantities  of  copper  contained  in  slags,  tailings  from  con- 
centration, and  similar  products. 

Heine's  modification  of  this  method,  as  described  by  Kerl,  is 
perhaps  the  most  convenient,  and  with  proper  solutions  for  com- 
parison, preserved  in  bottles  of  colorless  glass  and  of  exactly  the 
same  size,  yields  results  that  cannot  be  surpassed.  It  is  seldom 
employed  for  substances  containing  over  one  and  one-half  per  cent, 
of  copper,  and  may  be  relied  upon  to  show  differences  of  0.03  of 
one  per  cent. ;  results,  however,  depending  largely  upon  the  skill 
of  the  operator,  and  his  capacity  for  discriminating  almost  invisible 
shades  of  color. 


This  fire  assay  is  only  adapted  to  ores  free  from  sulphur  and 
other  metalloids.  Native  copper  is  the  principal  substance  dealt 
with,  though  oxides  of  copper  may  be  equally  well  determined  by 
this  method.  The  Lake  Superior  concentrates  consist  of  metallic 
copper,  and  sometimes  carry  up  to  50  per  cent,  of  titanic  iron  sand. 
Silica,  oxide  of  iron,  and  metallic  iron  from  the  stamp-heads  are 
also  usually  present. 

*  A  detailed  account, by  Mr.  M.  B.  Patch,  of  this  interesting  assay  was  given 
in  the  first  edition  of  this  work;  but  subsequently,  owing  to  the  rapid  accumu- 
lation of  material  that  is  of  more  value  to  the  majority  of  copper  metallurgists 
I  have  reluctantly  felt  obliged  to  omit  it. 



Sodium  bicarbonate,  borax,  potassium  bitartrate,  ferric  oxide, 
sand,  and  slag  from  the  same  operation  are  the  chief  fluxes  em- 

The  fluxed  sample  is  fused  in  a  wind  furnace  for  about  25  min- 
utes, the  resulting  button  being  almost  pure  copper,  and  its  weiglit 
agreeing  very  closely  with  battery  assays  of  the  same  sample. 

It  is  needless  to  say  that  very  much  depends  upon  the  skill  and 
experience  of  the  operator. 

The  same  assay  is  used  for  the  determination  of  copper  in  slags 
carrying  half  a  per  cent,  cf  copper,  and  less. 

I  append  a  table,  showing  the  fluxiug-formulas  for  various  sam- 
ples of  concentrates.  Also  tables  showing  results  of  this  assay  as 
compared  with  electrolysis. 

(These  tables  are  from  Mr.  Patch's  valuable  paper,  as  is  indeed, 
the  above  brief  abstract.) 



j  Borax. 







Iron  Ore. 


Per  cent. 



5  to  20 


j       60 
'       60 
1     100 
1     150 
!     190 





4. ! ! ! ! ! . 






The  results  obtained  by  this  method  are  surprisingly  accurate. 
Duplicate  determinations  of  the  lower  grade  samples  seldom  vary 
more  than  0.1  or  0.2.  A  difference  of  0.4  per  cent,  is  a  rare  oc- 
curence, even  in  the  higher  classes  of  mineral,  where  the  size  of  the 
metallic  fragments  renders  the  sampling,  and  even  the  weighing 
out,  of  a  correct  assay  a  matter  of  some  uncertainty. 

A  few  results  from  Mr.  Patch's  notes  will  confirm  these  state- 
ments. An  average  series  of  tests  on  cupola  slags  by  the  colori- 
metric  method  for  the  period  of  a  month,  duplicated  by  the 
fire  assay,  gave  a  result  0.05  per  cent,  lower  for  the  latter  test,  the 
slag  containing  about  0.5  of  one  per  cent. 

As  an  illustration  of  the  results  of  this  system  when  applied  to 
very  rich  ore,  a  comparative  test  was  made  for  eight  days  on  No.  1 
Calumet  &  Hecla  mineral,  with  the  following  results: 

Battery  assay 89. 100  per  cent. 

Fire  assay 88.819. 


A  similar  test  on  No.  2  Calumet  &  Hecla  mineral: 

Battery  assay , 77.590  per  cent. 

Fire  assay 77.657 

A  similar  test  with  various  samples: 

No.  Battery  Assay.  Fire  Assay. 

1 1  89.50"] 

I • VMean  =  89.544  g^-^J  lMean  =  89.92 

4[V.'  v.'.'.'.'.'.' '.'.'.'.'.'.'.  )  89^70  J 

5  1  77.401 

^ liMean  =  77.740  J!^-^^  U^ean  =  77.50 


J  77.40  J 

It  is  a  somewhat  curious  fact  that  the  slight  loss  of  about  0.25 
per  ceut.  of  copper,  which  results  from  the  passage  of  a  minute 
portion  of  the  metal  into  the  slag,  is  just  about  counter- 
balaiiced  by  the  impurities  in  the  copper  button  from  the  reduc- 
tion of  ferric  oxide,  the  amount  of  which  is  indicated  by  the 
following  analysis  of  copper  buttons — the  only  weighable  impurity 
being  iron: 

Copper.  Copper.  Copper. 

Per  cent.  Per  cent.  Per  cent. 

99.83  99.76  99.51 

99.84  99.80  99.87 
99.53                          99.46  99.79 

This  account  of  a  little  known  process  will  doubtless  remove  the 
impression  sometimes  held  by  chemists  that  the  Lake  Superior 
copper  assay  is  a  clumsy  and  imperfect  operation,  and  unworthy 
any  advanced  system  of  metallurgy. 



The  presence  of  a  large  proportion  of  copper  demands  special 
precautious  in  assaying  matte  or  bars  for  the  precious  metals. 

The  following  methods  will  be  found  simple  and  accurate,  and 
are  tliose  usually  employed  by  public  assayers,  and  at  the  principal 

We  notice  iu  the  outset  a  divergence  between  the  methods 
usually  employed  in  the  east  and  west  of  the  United  States.     Most 

*  This  description  is  taken  from  a  paper  read  by  Mr.  A.  H.  Ledoux  before 
tbe  American  Institute  of  Mining  Enjyineers,  October,  1894,  entitled  "A  Uni- 
form Method  for  the  Assay  of  Copper  Furnace  Materials  for  Gold  and  Silver."' 


of  the  Eastern  public  assayers,  as  well  as  those  employed  by  Eastern 
smelting  works,  use  what  may  be  called  a  wet  method,  but  is, 
strictly  speaking,  a  combination  method  of  assay.  While  there  are 
many  details  incidental  to  different  laboratories,  this  wet  method 
may  be  outlined  briefly  as  follows: 

For  Gold — One  assay-ton  of  the  copper-borings  or  matte  is  trans- 
ferred to  a  No.  5  beaker  with  a  clock-glass  cover.  The  sample  is 
treated  with  a  mixture  of  100  c. c.  of  water  and  oOc.c.  of  nitric  acid  of 
sp.  gr.  1.-42.  When  the  violent  action  has  ceased,  50  c.c.  more  of  the 
nitric  acid  is  added,  and  the  solution  is  gently  heated  until  everything 
soluble  has  been  dissolved.  The  contents  of  the  beaker  are  then  raised 
to  the  boiling  point,  the  cover  is  removed,  and  boiling  is  continued 
until  most  of  the  nitric  acid  has  been  expelled.  The  solution 
is  then  diluted  with  about  400  c.c.  of  water  free  from  chlorine,  5 
c.c.  of  concentrated  sulphuric  acid  is  added,  and  then  10  c.c.  of  a 
concentrated  solution  of  either  acetate  or  nitrate  of  lead.  The 
dense  white  precipitate  of  lead  sulphate  carries  down  with  it  the 
minute  particles  of  gold  which  may  be  suspended  in  the  solution. 

The  precipitate  is  then  allowed  to  settle  for  some  hours — over 
night,  if  possible.  It  is  then  filtered,  washed  once  or  twice  with 
water,  the  beaker  is  carefully  cleaned,  and  the  filter  and  contents, 
now  practically  free  from  copper,  are  partially  dried,  wrapped  in 
thin  lead-foil,  and  transferred  to  scorifiers;  enough  test-lead  is 
added  to  bring  the  total  lead  present  u^i  to  50  grams,  a  pinch  of 
borax  glass  is  placed  on  top,  and  the  scorification  is  conducted  as 
usual.  It  is  necessary  to  raise  the  temperature  gradually  until  the 
paper  has  been  consumed  and  the  contents  of  the  scorifier  melted 
down.     Cupellation  is  conducted  in  the  usual  manner. 

This  method  is  intended  for  the  determination  of  gold;  but 
enough  silver  will  be  present  to  allow  the  bead  to  be  parted. 
When,  however,  considerable  gold,  say  two  or  three  ounces  per  ton 
(O.Olf^,)  is  supposed  to  be  present,  it  is  well  to  add  a  drop  of  salt 
solution  to  the  original  nitric  acid  solution,  to  precipitate  some 
of  the  silver  along  with  the  lead,  or  else  to  add  a  small  amount  of 
pure  silver  at  the  time  of  scorification.  It  is  important  not  to 
precipitate  all  the  silver,  as  in  that  case  there  might  be  an  excess 
of  salt  which  might  liberate  chlorine  and  vitiate  the  results  as  to 

For  Silver. — The  usual  method  employed  in  the  East  for  the 
assay  of  copper  bars,  mattes,  ores,  etc.,  containing  silver  is  like- 
wise modified  in  different  laboratories.   These  modifications  vary. 


as  a  rule,  witli  the  supposed  richness  in  silver  of  the  sample 
treated  The  sample  is  dissolved  in  dilute  nitric  acid,  as  described 
in  the  above  method  for  gold.  To  the  solution,  after  the  addition 
of  sulphuric  acid  and  before  that  of  lead  acetate,  a  solution  of 
chloride  of  sodium  is  added  in  a  sufficient  quantity  to  throw  down 
all  the  silver,  the  addition  being  gradual,  and  avoiding  a  great* 
excess  (as  silver  chloride  is  more  or  less  soluble  in  sodium  chloride 
solution);  then  the  lead  acetate  is  added,  the  solution  is  well  stirred, 
and  the  mixed  precipitate  of  lead  sulphate  and  silver  chloride  is 
allowed  to  settle  as  in  the  gold  determination.  The  rest  of  the 
jirocess  is  conducted  exactly  as  in  the  previous  case  for  gold. 
Where  any  considerable  amount  of  gold  is  present  it  is  of  course 
necessary  to  part  the  beads  and  deduct  the  weight  of  gold  present, 
which  otherwise  would  be  weighed  as  silver,  thus  erroneously 
increasing  the  proportion  of  this  metal.  The  gold  obtained  by  this 
parting  is  usually  less  than  the  figures  obtained  by  the  special  assay 
for  gold,  because  some  of  the  gold  is  dissolved  by  chlorine  through 
the  excess  of  sodium  chloride  employed. 

Some  assayers  determine  the  gold  and  silver  at  one  operation  by 
taking  the  filtrates  from  the  gold  and  lead  sulphate  precipitate 
obtained  as  above  described,  precipitating  the  silver  in  this  solution 
fis  chloride,  adding  more  lead  acetate,  and  after  filtering,  combin- 
ing the  two  filter  papers,  one  containing  the  gold  and  the  other  the 
silver,  and  uniting  them  for  one  scorification  and  subsequent 
■cupellacion.  Tliis  method  is  more  economical  for  the  assayer,  and 
has  the  advantage  also  of  two  filtrations  for  gold,  catching  any  fine 
particles  which  might  pass  through  the  first  filter;  but  on  the  other 
hand  it  takes  more  time,  because  the  same  solution  is  twice  settled. 
In  the  first  method,  the  settling  of  tlie  gold  and  silver  precipitates 
goes  on  simultaneously. 

In  the  West,  the  all-fire  method  is  employed  almost  exclusively, 
so  far  as  I  can  ascertain.  In  the  Omaha  and  Grant  works,  for 
example,  ten  portions  of  sample,  of  one-tenth  A.  T.  each,  are 
weighed  out  and  scorified  with  50  grams  of  test-lead,  one-half  of 
which  lead  is  mixed  with  the  sample  and  the  remainder  used  to 
cover  it  in  the  scorifier.  One  gram  of  borax  is  added.  The 
lead  buttons  obtained  by  the  scorification  are  cupelled  separately, 
but  the  ten  beads  are  weighed  together.  The  cupels  are  then, 
ground  np  and  fused  in  five  lots  of  two  each,  with  the  following 
charge:  Litharge,  90;  soda,  50;  borax-glass,  50,  and  argols,  3 
grams.     The  five  buttons  are  cupelled  and  the  silver  is  added  to  that 


obtained  in  the  first  operation,  representing  the  loss  in  scorification. 
All  the  beads  are  then  parted  for  gold,  which  is  deducted  from  the 
total  weight  as  usual. 

My  experience  shows  that  the  determination  of  gold  obtained  by 
this  process  is  usually  higher  than  where  the  wet  process  previously 
described  is  employed.  It  may  be  well  to  give  certain  instances 
in  my  own  experience.  On  high-grade  copper  bullion,  which 
contains  on  an  average  about  400  ounces  of  silver  per  ton  (1.37^), 
the  results  were: 

Fire  Assay.  Wet  and  Dry  Assay. 

Gold,  ounces  per  ton 1.06  [0.00364  per  cent.]        0.92  [0.00316  per  cent.] 

"       1.32  [0.00458  per  cent.]         1.24  [0.00426  per  cent.] 

"      0.34  [0.00117  per  cent.]         0.20  [0.00069  per  cent.] 

In   bullion  containing  300  ounces  of  silver  per  ton  (1.03^): 

Fire  Assay.  Wet  and  Dry  Assay. 

Gold,  ounces  per  ton 4.06  [0.01395  per  cent.]        3.96  [0.0136    per  cent.] 

2.76  [0.00948  per  cent.]        2.56  [0.00879  per  cent.] 

2.72  [0.00934  per  cent.]        2.44  [0.00838  per  cent.] 

In  matte  containing  60  per  cent,  of  copper  and  60  ounces  of 

Fire  Assay.  Wet  and  Dry  Assay. 

Gold,  ounces  per  ton 0.24  [0.00082  per  cent.]        0.20  [0.00069  per  cent.] 

The  two  processes  usually  agree  very  closely  for  silver,  provided  the 
cupel-absorption  is  determined  when  the  silver  is  assayed  by  the 
combination  wet  process.  This  cupel-absorption  is  very  much  less 
by  the  wet  process  than  by  the  all-fire  method,  because  by  the 
former  the  copper  has  been  eliminated,  and  is  not  present  to  help 
carry  the  silver  into  the  cupel.  In  some  instances,  where  sub- 
stances are  present  which  would  cause  volatilization  of  silver  in 
scorification,  the  wet  assay  gives  higher  figures,  because  the  inter- 
fering substance  has  been  removed  by  the  acid. 

The  Western  all-fire  process  for  mattes  is  similar  to  that  employed 
for  bars,  except  that  a  second  scorification  is  sometimes  necessary 
before  cnpellation.  The  second  scorification  is  usually  performed 
iu  a  small  2:^-inch  scorifier,  enough  test-lead  being  added  to  the 
button  obtained  from  the  first  scorification  to  make  the  lead  present 
not  less  than  35  grams. 

The  above  descriptions,  as  will  readily  be  seen,  are  in  the  baldest 


outline;  and  it  must  not  be  inferred  by  those  iuturested  that  all 
precautious  are  not  adopted  to  make  the  results  correct;  such,  for 
instance,  as  igniting  and  dissolving  any  sulphur-balls  which  may 
form  when  the  matte  or  sulphuret-ores  are  dissolved  in  acid,  and 
adding  the  product  to  the  main  solution  before  precipitating  the 
silver  with  lead.  This  precaution  is  hardly  necessary,  however,  as 
the  very  small  amount  of  matte  or  ore  held  by  the  sulphur  would 
bo  decomposed  in  the  scorification. 

Each  of  these  methods  in  the  hands  of  assayers  skilled  in  its 
application  will  produce  very  "uniform  results;  and  yet,  as  will  be 
seen  from  the  few  comparisons  given  above,  any  assayer  running 
the  two,  side  by  side,  will  get  divergent  figures  for  gold. 


The  following  method  for  the  estimation  of  sulphur  in  materials 
containing  it  in  the  form  of  sulphides  not  decomposable  by 
hydrochloric  acid  is  found,  in  practice,  to  be  exceedingly  accurate 
and  convenient.  The  great  majority  of  methods  now  practised 
consists  in  oxidizing — either  in  the  dry  way  or  wet  way — the 
sulphur  to  sulphuric  acid,  and  estimating  the  latter  gravimetri- 
cally  or  by  titration. 

The  method  about  to  be  described  is  likewise  based  on  this 
principle,  and  is  a  combination  of  well-known  reactions.  What  is 
claimed  for  it  is  that  it  is  especially  adapted  to  the  quick  deter- 
mination of  the  sulphur  in  roasted  copper  ores  and  cupriferous 

The  conversion  of  the  sulphur  into  an  alkaline  sulphate  is 
effected  by  fusion  with  potassium  hydrate  and  sodium  peroxide, 
and  the  amount  of  sulphur  is  then  ascertained  in  the  usual  manner 
— gravimetrically,  when  accuracy  is  the  principal  object — by  the 
Wildenstein  method,  when  rapidity  is  aimed  at. 

The  details  of  the  process  are  as  follows:  Five  to  six  grams  of 
caustic  potash  (pure  by  alcohol)  are  fused  in  a  nickel  crucible  and 
heated  until  the  excess  of  water  is  expelled.  The  size  of  the  flame 
is  now  reduced  so  that  the  contents  of  the  crucible  just  remain 
liquid,  and  0.5  grams  of  the  finely  powdered  material  introduced 
in  small  portions.  A  gram  of  sodium  peroxide  is  then  added  while 
the  heat  is  gradually  increased  to  redness,  and  this  is  maintained 
for  a  few  minutes.     After  cooling,  the  fused  mass  is  dissolved   in 

*Tbis  method  was  devised  by  Harry  F.  Keller  and  Philip  Maas,  and  com- 
municated to  the  Franklin  Institute,  January  18,  1895. 


water  and    the   solution   filtered  with    the  aid  of  a  piiinp.     The 
undissolved  residue  is  washed  four  or  five  times  with  hot  water. 

The  colorless  filtrate  is  acidified  with  hydrochloric  acid  (8  to  9 
c.c,  sp.  gr.  1.2)  and  boiled  to  expel  carbonic  acid.  (Before 
filtering,  the  hquid  is  often  colored  purple  by  a  small  amount  of 
ferrate  of  potassium;  a  blue  color  in  the  filtrate  indicates  that  too 
much  potash  was  used.) 

If  the  estimation  is  to  be  made  gravimetrically,  the  sulphate  of 
barium  is  precipitated  from  the  boiling  liquid  in  the  usual 
manner.  In  case,  however,  titration  is  resorted  to,  the  liquid 
is  made  alkaline  with  ammonia  (about  .5  c.c,  sp.  gr.  0.9).  A 
slight  excess  of  barium  chloride  solution  is  added  from  a  burette, 
and  the  excess  measured  with  an  equivalent  solution  of  bichromate 
of  potassium.  A  distinct  yellow  color  of  the  liquid  marks  the 
end  of  the  reaction.  After  a  little  practice  it  is  generally  easy  to 
strike  this  point,  thongli  it  will  sometime  happens  that  the  precipi- 
tate does  not  settle  rapidly.  In  such  doubtful  cases  portions  of 
the  liquid  should  be  filtered  oif.  Care  should  also  be  taken  that 
the  liquid  does  not  become  too  dilute.  It  is  convenient  to  prepare 
the  standard  solution  of  such  strength  that  1  c.c.  equals  0.005  grams 
of  sulphur,  i.e.^  indicates  1  per  cent,  in  a  sample  weighing 
0.5  grams. 

The  solution  of  barium  chloride  is  prepared  by  dissolving  38.109 
grams  of  the  crystallized  salt  to  a  liter,  while  the  bichromate  solu- 
tion should  contain  23  grams  of  the  salt  per  liter. 

To  test  the  accuracy  of  this  method  a  considerable  number  of 
determinations  were  made  of  the  sulphur  in  a  typical  roasted  copper 
ore  from  Montana. 

By  oxidation  with  nitric  acid  and  with  aqua  regia  the  percentage 
of  sulphur  in  this  material  had  been  found  to  be  7.095  per  cent, 
and  T.l-t  per  cent,  respectively. 

Somewhat  lower  results  were  obtained  by  fusion  with  caustic 
potash  and  potassium  chlorate,  a  method  which  had  been  used  by 
one  of  us  to  control  the  workings  of  a  lead-ore  roasting  furnace. 
The  figures  varied  from  G.T8  per  cent,  to  G.92  per  cent. 

Our  first  attempts  to  oxidize  the  ore  by  means  of  sodium  peroxide 
were  not  successful.  By  using  10  grams  of  potash  and  3  to  5 
grams  of  peroxide,  figures  much  lower  than  those  given  before 
resulted.  The  oxidation  was  evidently  incomplete.  When  bromine 
water  was  added  to  the  solution  of  the  fused  mass,  6.82  per  cent, 
of  sulphur  were  obtained. 


To  our  surprise,  a  higher  percentage  was  also  found  when  less  of 
the  peroxide  was  employed.  Thus  with  10  grams  of  potash  and  1 
gram  of  peroxide,  the  determiuations  averaged  6.8  per  cent.  The 
large  excess  of  alkali  employed  in  these  fusions  invariably  caused 
the  solution  of  some  copper,  which  renders  titration  impossible. 
Our  next  step,  therefore,  was  to  reduce  the  amount  of  potash. 

When  5  grams  of  hydrate  and  1  gram  of  peroxide  were  taken, 
the  filtered  solution  of  the  fused  mass  was  entirely  free  from  the 
blue  tint  produced  by  the  copper,  and  it  is  seen  from  the  following 
figures  that  the  oxidation  of  the  sulphur  was  complete: 

1 6.71  per  cent,  sulphur. 

3 6.82 

3 6.79 

Average 6.77         "  " 

Volumetric  estimations  gave  the  following  results: 

1 6.74  per  cent,  sulphur. 

3..... 6.86 

3 6.89 

4 7.14 

5 6.97 

Average 6.92        " 

Another  series   of  determinations,  in  which   a  preparation   of 
potash  marked ptiriss pro  analys*  was  used,  yielded: 

1 6.70  per  cent,  sulphur. 

3 7.09 

3 ....6.71 

4 6.85        "  " 

5 6.79        "  " 

6 7.00        "  " 

Average 6.85        "  ** 

A  final  series,  in  which  the  dircetions  given  in  this  paper  were 
strictly  adhered  to,  resulted  as  follows: 

*A  correction  of  0.35  per  cent,  was  necessary  in  this  case,  the  potash  being 
less  free  from  sulphur  than  that  labeleJ  "  pure  by  alcohol." 



1 6.9  per  cent,  sulphur. 

2 6.75 

3 7.15 

4 7.05 

5 7.05 

6 7.10 

7 7.14 



The  time  required  for   the  volumetric   assay   does  not  exceed 
thirty  minutes. 



*RoASTiisrG  or  calcination,  used  indiscriminately  in  the  language 
of  the  American  copper  smelter,  signifies  the  exposure  of  ores  of 
metals  containing  sulphur,  arsenic,  and  other  metalloids,  to  a 
comparatively  moderate  temperature,  with  the  purpose  of  ejecting 
certain  chemical,  and  rarely  mechanical,  changes  required  for 
their  subsequent  treatment.  This  definition  is  restricted  to  the 
dry  metallurgy  of  copper,  and  does  not  take  into  consideration 
chloridizing  roasting,  roasting  with  sulphate  of  soda,  and  other 
well-known  variations,  which  belong  either  to  the  metallurgy  of 
the  precious  metals  or  to  the  wet  treatment  of  copper  ores. 

The  care  and  attention  which  should  be  devoted  to  this  prepara- 
tory process  cannot  be  too  strongly  insisted  on,  nor  can  any  one 
carry  out  either  this  apparently  simple  roasting  or  the  following 
fusion  to  the  best  advantage,  who  is  not  thoroughly  familiar  with 
the  striking  chemical  cbanges  that  in  every  calcination  follow  closely 
upon  each  other,  and  by  which  the  sulphides  and  arsenides  of  the 
metals  are  transformed  at  will  into  a  succession  of  subsulphides,  sul- 
phates, subsulphates,  and  oxides.  These,  reacting  upon  each  other 
according  to  fixed  and  well-known  laws,  enable  the  metallurgist 
at  his  pleasure  to  produce  every  grade  of  metal  from  black  copper 
to  a  low-grade  matte  that  shall  contain  nearly  all  the  metallic  con- 
tents of  the  ore  in  combination  with  sulphur.  To  avoid  constant 
repetition,  it  may  be  understood  that  in  speaking  of  calcination, 
when  sulphur  is  mentioned,  its  more  or  less  constant  satellites, 
arsenic  and  antimony,  are  also  included,  their  behavior  being 
somewhat  similar  under  ordinary  circumstances.  These  very 
different  products,  as  well  as  the  amount  of  ferrous  oxide,  the  most 
important  basic  element  of  every  copper  slag,  result  solely  from 

*  In  English  metallurgical  literature,  the  term  roasting  is  applied  exclusively 
to  that  process  in  which  copper  matte  in  large  fragments  is  exposed  on  the 
hearth  of  a  reverberatory  furnace  to  an  oxidizing  atmosphere,  and  a  moderate 
but  gradually  increasing.temperature. 

76  MODEKX    CUri^EK   bAlELTlXG. 

the  degree  to  which  the  calciuation  is  carried.  In  fact,  it  mav  he 
takeu  as  literally  true,  that  the  composition  of  both  the  valuable 
and  waste  products  of  the  fusion  of  any  sulphide  ore  of  copper  is 
determined  irrevocably  and  entirely  in  the  roasting-furnace  or 
stall.  A  more  thorough  study  of  the  reactions  just  referred  to  will 
be  found  in  its  proper  place.  Enough  has  here  been  said  not  only 
to  exphiin  the  author's  object  in  devoting  so  much  attention 
to  this  process,  but  also  to  induce  such  smelters  as  are  not 
already  thoroughly  familiar  with  the  theory  of  calcination  to 
endeavor  to  become  so  if  they  desire  to  ever  excel  in  the  economical 
treatment  of  sulphide  ores. 

The  varieties  of  calcination,  as  applied  to  the  dry  treatment  of 
copper  ores,  are  at  most  two: 

1  The  oxidizing  roasting,  which  is  necessarily  combined  with 

"2.  The  reducing  roasting,  limited  in  its  application  almost 
exclusively  to  substances  containing  much  antimony  or  arsenic. 

Plattuor's  admirable  work  on  Rostproresse  contains  the  whole 
theoretical  part  of  calcination;  but  a  foreign  language  is  a  barrier 
to  many  ardent  students  of  metallurgy,  and  his  descriptions  and 
plans  of  furnaces  and  apparatus  apply  to  those  in  use  during  the 
past  generation.  A  modern  treatise  on  roasting,  regarding  the 
subject  principally  from  a  practical  standpoint,  and  adapted  to 
present  American  conditions,  seems  desirable.  Such  a  treatise, 
however,  could  not  attain  the  highest  degree  of  usefulness  without 
a  consideration  of  the  theory  of  calciuation  sufficient  to  enable  and 
encourage  all  who  make  use  of  the  more  practical  part  to  follow 
with  ease  the  chemical  reactions  on  which  the  process  is  based. 

A  sufficient  idea  of  the  chemical  reactions  that  occur  in  this 
important  metallurgical  process  may  be  obtained  by  following  an 
ordinary  pyritous  ore  in  its  passage  through  the  roasting-furnace, 
and  carefully  noting  all  the  changes  that  it  undergoes  from  the 
moment  of  its  introduction  until  it  is  ready  for  the  succeeding 
fusion;  nor  are  the  conditions  in  either  roast-heaps  or  stalls  so 
different  as  to  require  any  separate  consideration. 

A  typical  ore  for  this  purpose  might  consist  of  a  large  proportion 
of  pyrite,  say  45  per  cent.,  some  20  per  cent,  of  chalcopyrite 
(containing  about  one-tiiird  copper),  with  a  slight  admixture  of 
zinc-blende,  galena,  and  sulphide  of  silver,  while  the  remainder 
of  the  ore  would  usually  consist  of  quartz  or  silicious  material, 
which  may  be  regarded  as  practically  inert  in  its  effect  upon  the 


process  of  calcination.  A  charge  of  such  ore,  being  introduced 
upon  the  hearth  of  a  roasting-furnace  still  at  a  bright  red  heat 
from  the  preceding  operation,  exerts  a  powerfully  cooling  influ- 
ence upon  the  glowing  brick-work,  and  within  ten  or  fifteen 
minutes  reduces  the  temperature  to  a  point  below  the  ignition 
point  of  sulphur,  the  ore  at  the  same  time  giving  off  its  moisture, 
and  gaining  so  much  heat  that  a  very  slight  aid  from  the  fuel  on 
the  grate  is  sufficient  to  start  the  oxidation  of  the  iron  pyrites,  as 
shown  by  the  blue,  flickering  flame  that  plays  over  the  surface  of 
the  charge,  beginning  at  that  portion  of  the  same  that  borders  on 
the  already  hot  charge  occupying  the  adjoining  hearth,  and 
gradually  advancing  toward  the  rear,  until  every  square  inch  of 
surface  is  in  a  state  of  active  combustion.  The  rapidity  of  this 
process  of  oxidation  varies  according  to  the  degree  of  temperature 
and  the  sharpness  of  the  draught,  but  should  not  occupy  more 
than  an  hour  from  the  first  introduction  of  the  charge.  The 
composition  of  iron  pyrites  (FeS.,)  is  such  that,  while  one  atom  of 
sulphur  is  united  to  the  iron  with  considerable  tenacity,  the 
second  atom  is  held  by  very  feeble  bonds,  and  becoming  volatile 
at  the  moderate  temperature  of  the  calcining  furnace,  unites  with 
the  oxygen  of  the  air,  forming  sulphurous  acid  (SOg),  which 
escapes  in  the  form  of  an  invisible  gas.  This  reaction  is  accom- 
panied by  a  very  considerable  evolution  of  heat  and  the  flickering 
blue  flame  already  mentioned.  Being  entirely  dependent  upon 
the  oxygen  derived  from  the  air,  this  reaction  is  confined  prin- 
cipally to  the  surface  of  the  charge,  which,  if  left  undisturbed,, 
would  soon  undergo  a  slight  fusion,  causing  a  caking  of  the  ore, 
and  still  further  hindering  the  extension  of  the  process.  It  is 
therefore  Just  at  this  point  that  the  necessity  for  frequent  and 
vigorous  stirring  becomes  strikingly  apparent.  By  this  manipu- 
lation, any  incipient  crust  that  may  have  formed  is  broken  up,  the 
temperature  of  the  layer  of  ore  is  equalized  throughout  its  entire 
depth,  and  fresh  particles  of  ore  are  constantly  exposed  to  the 
influence  of  the  air. 

The  stirring  should  begin  on  the  first  appearance  of  the  blue 
flame,  and  continue  for  ten  minutes  at  a  time,  with  equal  intervals; 
of  rest,  during  which  time  the  working  openings  should  be  closed, 
while  an  ample  air  supply  is  admitted  through  the  regular 
channels  provided  for  this  purpose.  The  stirring  should  take 
place  from  both  sides  of  the  furnace  at  the  same  time,  and  should. 


be  systematic,  vigorous.and  thorough;  extending  to  the  very  bot- 
tom of  the  charge,  and  omitting  no  portion  of  the  ore. 

During  this  period  of  roasting,  and  until  the  disappearance  of 
the  blue  flame,  the  roast  gases  consist  almost  exclusively  of  sul- 
phurous acid,  together  with  steam  from  the  moisture  present,  and 
the  invariable  products  of  the  combustion  of  the  fuel. 

It  will,  of  course,  be  understood  that  tlie  SO2,  and  other  roast 
gases, form  but  a  small  proportion — seldom  more  than  2  per  cent. 
— of  the  air  issuing  from  a  calciner  stack;  atmospheric  air  always 
being  present  in  overwhelming  proportions.  The  SOg  results 
from  the  direct  oxidation  of  one  atom  of  the  sulphur  contents  of 
the  iron  pyrites,  or,  when  the  temperature  is  somewhat  high,  of 
the  absolute  volatilization  of  this  atom  of  sulphur  as  sulphur,  and 
its  immediate  combustion  to  SO2. 

The  next  stage  of  the  process  may  be  reckoned  from  tlie  begin- 
ning of  the  oxidation  of  the  iron  of  the  pyrites,  and  also  of  its 
second  atom  of  sulphur.  This  is  a  much  less  rapid  and  vigorous 
process  than  the  preceding,  and  is  attended  by  the  formation  of  a 
certain  amount  of  sulphuric  acid,  in  addition  to  the  sulphurous 
acid,  which  is  still  generated  in  large  quantities.  The  means  by 
which  the  former  acid  was  produced  was  not  clearly  understood 
until  Plattner's  patient  and  ingenious  researches  developed  the 
"contact  theory,"  according  to  which  sulphurous  acid  and  the 
oxygen  of  the  air,  in  the  presence  of  large  quantities  of  heated 
quartz,  or  other  neutral  material,  combine  to  form  sulphuric  acid, 
which  may  escape  invisible,  or  in  the  form  of  white  vapors  when 
hydrated,  or  may  in  the  instant  of  its  formation  combine  with  any 
strong  base  that  may  be  present. 

In  the  case  under  consideration,  protoxitle  of  iron  (FeO),  arising 
perhaps  from  the  very  particle  of  pyrites  whose  oxidation  gave  rise 
to  the  sulphuric  acid,  is  at  hand;  and  while  the  greater  proportion 
of  the  sulphuric  acid  formed  escapes  into  the  atmosphere,  a  certain 
amount  combines  with  the  protoxide  of  iron  to  form  ferrous  sul- 
phate, whose  presence  may  easily  be  detected,  owing  to  its  solu- 
bility in  water. 

From  the  very  commencement  of  the  formation  of  sulphuric 
acid,  a  new  and  powerful  oxidizing  agent  is  gained,  as  the  protosul- 
phate  of  iron  is  easily  broken  up  by  heat.  The  decomposition  of 
its  acid  into  SO2  and  0  promotes  the  oxidation  of  other  sulphides 
present  to  sulphates,  while  the  protoxide  of  iron  is  raised  to  the 
sesquioxide   of  that  metal — a  tolerably  stable  compound,  and  one 


usually  found  in  large  quantities  in  thoroughly  roasted  pyritic 
ores.  Before  the  complete  decomposition  of  the  ferrous  sulphate 
has  occurred,  and  indeed  while  some  considerable  proportion  of 
sulphide  of  iron  may  yet  remain,  an  analogous  process  takes  place 
with  the  chalcopyrite,  its  ferruginous  portion  following  almost 
precisely  the  same  course  as  the  iron  pyrites,  while  its  copper  con- 
tents are  transformed  into  cupric  sulpiiate,  which,  on  the  addition 
of  water,  becomes  copper  vitriol,  easily  recognized  by  its  color  and 
by  several  simple  and  well-known  tests. 

As  the  process  continues,  and  the  temperature  is  gradually 
raised,  this  salt  also  undergoes  decomposition,  yielding  at  first  a 
basic  sulphate  of  copper,  which,  upon  losing  its  acid,  becomes  a 
dioxide  and  eventually  a  protoxide  of  that  metal.  These  last 
ohauges,  however,  require  a  protracted  high  temperature. 

The  oxidation  of  the  iron  present  is  pretty  well  advanced  at  the 
time  of  the  maximum  formation  of  cupric  sulphate;  but  it  is  not 
until  the  decomposition  of  at  least  75  per  cent,  of  the  last-named 
salt  that  the  formation  of  sulphate  of  silver  begins  with  any  con- 
siderable energy.  When  once  fairly  started,  however,  this  interest- 
ing and  important  reaction  progresses  with  great  rapidity,  and  the 
decomposition  of  the  comparatively  large  proportion  of  sulphate  of 
oopper  present  furnishes  ample  oxidizing  influence  for  the  minute 
■quantities  of  sulphide  of  silver.  The  maximum  formation  of 
the  latter  substance  usually  coiucides  with  the  almost  entire 
destruction  of  the  former  salt,  and  it  is  at  this  point  that  the 
Ziervogel  calcination  should  terminate,  as  any  further  exposure 
of  the  silver  salt  to  heat  lessens  its  solubility  in  water,  and  may 
■even  threaten  its  existence.  The  complete  decomposition  of  the 
argentic  sulphate  is  only  accomplished  by  a  long  exposure  to  a 
high  temperature,  which  is  now  easily  borne  by  most  ores  and 
mattes,  the  easily  melted  sulphides  having  been  converted  into 
almost  infusible  oxides  and  basic  sulphates. 

Galena  (sulphide  of  lead),  when  present,  is  converted  almost 
entirely  into  a  sulphate  of  that  metal,  which,  by  a  higher  tempera- 
ture, is  partially  decomposed  with  the  evolution  of  sulphurous  acid 
and  the  final  production  of  a  mixture  of  free  oxide  of  lead  with 
sulphate,  the  proportions  of  these  two  substances  varying  accord- 
ing to  the  quantity  of  foreign  sulphides  present. 

Zinc-blende  requires  a  higher  heat  for  its  thorough  oxidation 
than  any  of  the  preceding  sulphides,  but  with  care  may  be 
eventually  changed  into  an  oxide,  although  a  certain  amount  of 


basic  sulphate  of  ziuc  nearly  always  remains.  This  includes  all 
the  sulphides  assumed  to  have  been  present  in  the  ore  under  con- 
sideration, nor  will  others  be  encountered  in  practice  unless  under 
very  exceptional  circumstances.  Sulphide  of  manganese  is  an 
occasional  unimportant  constituent  of  mattes,  and  presents  no 
particular  difficulty  in  calcining,  being  easily  oxidized  to  a  basic 
sulphate,  insoluble  in  water,  which  is  stable  except  at  the  highest 
roasting  temperatures,  when  it  yields  up  its  acid  in  the  shape  of 
SOi,  and  remains  as  a  mixture  of  maugauous  and  manganic  oxides. 

The  gangue-rock  of  copper  ores,  beiug  usually  silicious,  under- 
goes no  change  and  exerts  no  influence  upon  the  calcining  process, 
except  in  so  far  as  it  assists  in  the  oxidation  of  sulphurous  to 
sulphuric  acid  by  contact,  as  already  mentioned. 

Calc-spar  loses  its  carbonic  acid  and  is  converted  into  gypsum 
(calcium  sulphate),  while  heavy  spar — sulphate  of  baryta — under- 
goes no  change,  except  in  the  presence  of  a  powerlul  reducing 
atmosphere  and  at  a  high  temperature,  when  it  may  be  changed 
into  sulphide  of  barium.  This  is  soluble  in  water,  and  it  has  been 
suggested  to  use  its  solubility  to  remove  it  when  its  presence  is 
particularly  objectionable.  A  number  of  trials  in  this  direction 
were  made  by  the  author  in  1872  on  the  heavy  spar  ores  of  Mount 
Lincoln,  Colorado,  with  very  poor  results;  as  it  was  found 
extremely  difficult  to  reduce  the  barium  sulphate  to  sulphide  with- 
out mixing  an  amount  of  coal-dust  with  the  ore  at  least  equal  to 
the  weight  of  the  heavy  spar  present — from  30  to  40  per  cent. — 
while  the  BaS  formed  at  this  high  temperature  is  only  partially 
soluble  in  water. 

Arsenic  and  antimony,  when  present,  are  usually  combined  with 
some  metallic  base,  and  behave  like  sulphur  to  a  certain  extent; 
but  they  give  off  a  much  smaller  proportion  as  volatile  antimonious 
and  arsenious  acids,  while  they  combine  to  a  much  greater  extent 
with  the  metallic  bases,  forming  salts  difficult  to  decompose  and 
extremely  injurious  to  the  quality  of  the  copper. 

Under  such  circumstances  the  roasting  should  be  continued  in 
the  usual  manner  until  all  the  sulphides  present  are  oxidized  and 
the  resulting  sulfihates  for  the  most  part  decomposed.  At  this 
stage,  from  4  to  0  per  cent,  of  charcoal  dust.or  fine  bituminous  or 
anthracite  coal-screenings, should  be  thrown  upon  the  charge  and 
thoroughly  incorporated  with  it  by  vigorous  stirring,  the  heat  at 
the  same  time  being  raised  to  the  highest  practicable  limits. 
The  antimonates  and    arsenates   of   iron  and  copper  are  rapidly 


reduced  by  this  means,  and  a  considerable  projiortion  of  the 
injurious  metalloids  is  volatilized,  much  to  the  benefit  of  the 
resulting  copper.  The  charge  should  remain  in  the  furnace 
until  all  the  incorporated  carbon  is  consumed. 

In  the  foregoing  description  the  process  of  calcination  has  been 
carried  much  further  than  is  generally  needed,  or  even  desired,  in 
an  ordinary  oxidizing-roasting  as  a  preliminary  to  fusion. 

Sufficient  sulphur  must  always  be  present  in  the  smelting  mix- 
ture to  prevent  the  formation  of  too  rich  a  matte,  which  entails 
heavy  losses  in  metal,  and  other  injurious  consequences.  But  it  is 
not  a  simple  matter  to  determine  in  advance  exactly  the  amount  of 
sulphur  necessary  to  produce  a  matte  of  any  given  grade.  This 
depends  not  only  upon  the  cliaracter  of  the  furnace  process  to  be 
employed — that  is,  whether  blast  or  reverberatory — but  also  to  a 
considerable  extent  upon  the  manner  in  which  the  residual  sulphur 
is  combined  with  the  bases  present;  the  rapidity  of  the  fusion;  the 
quality  of  the  fuel;  tlie  volume  and  pressure  of  the  blast;  the 
character  of  the  gangue  and  flux;  and  numerous  other  factors. 
Whatever  may  be  the  condition  of  affairs,  however,  it  may  be 
pretty  safely  predicted  that  the  percentage  of  the  resulting  matte 
in  copper  will  almost  invarial)ly  be  very  considerably  lower  than  is 
either  expected  or  desired,  so  that  there  is  little  danger  that  the 
calcining  department  of  any  newly  constructed  plan  will  have  too 
great  a  capacity  in  proportion  to  the  rest  of  tue  establishment,  and 
many  serious  errors  and  disappointments  can  be  traced  directly  to 
this  habit  of  over-estimating  the  probable  quality  of  the  matte 
and  failing  to  provide  sutficient  calcining  appliances. 

In  case  of  calcination  previous  to  smelting  in  reverberatories,  it 
is  well  to  avoid  an  excess  of  air  toward  the  close  of  the  roasting 
process — a  precaution  easily  effected  by  closing  the  working 
openings  as  far  as  possible,  the  rabble  passing  through  a  hole  in 
the  center  of  a  divided  door,  while  the  passage  of  any  considerable 
proportion  of  undecomposed  air  through  the  grate  is  rendered 
unlikely  by  the  lively  fire  that  belongs  to  this  period.  By  these 
precautions  the  oxidation  of  any  large  proportion  of  the  iron 
present  to  a  sesquioxide  is  prevented,  the  latter  being  infusible 
and  unfit  to  enter  the  slag  until  it  is  reduced  to  a  protoxide. 
This  reduction  takes  places  instantaneously  in  the  powerful 
carbonic-oxide  atmosphere  that  prevails  in  the  blast  furnace;  but 
in   the  almost  neutral   atmosphere   of  the  ordinary  reverberatory 


the  sulphur  alone  plays  the  part  of  a  reducing  agent,  and  a 
charge  composed  of  the  sesquioxide  of  iron  will  be  found  materially 
to  delay  the  process  of  fusion,  besides  producing  a  thick  and  foul 
scoria.  The  natural  remedy  is  the  admixture  of  a  few  per  cent,  of 
fine  coal  stirred  thoroughly  into  the  mass  of  the  ore,  and  fired 
on  vigorously. 

Some  kind  of  an  idea  may  be  obtained  of  the  probable  composi- 
tion of  the  matte  to  be  produced  at  any  given  time  by  the  ordinary 
''matte  fusion  assay,"  as  given  in  all  works  on  assaying,  wherein 
the  ore  to  be  tested  is  rapidly  melted  with  merely  enough  borax 
and  silicious  flux — say,  100  per  cent,  of  borax  and  an  equal  amount 
of  pulverized  window-glass — to  flux  its  earthy  constitutents,  some 
10  per  cent,  of  argols,or  other  reducing  agents, being  also  added. 

But  the  results  are  far  from  satisfactory,  and  after  patiently 
using  it  for  some  two  years,  and  being  oftener  misled  than  guided 
by  its  results,  I  discarded  it  completely,  and  trusted  principally  to 
the  eye,  occasionally  aided  by  the  following  calculation,  which  gives 
better  results  than  any  other  familiar  to  me: 

Taking  the  contents  of  copper  in  the  charge  as  a  standard  for 
comparison,  sufficient  sulphur  should  be  allotted  to  it  to  form  a 
subsulphide,  the  excess  of  sulphur  still  remaining  being  supplied 
with  sufficient  iron  to  form  a  monosulphideof  that  metal.  If  other 
metals  are  present,  such  as  lead,  zinc,  or  manganese,  three-fourths 
of  the  former,  one-half  of  the  second,  or  one-fourth  of  the  latter 
substance  may  be  first  considered  as  forming  a  monosulphide  with 
the  sulphur,  there  being  in  such  a  case  much  less  of  the  metalloid 
left  to  take  up  iron.  This  rule  gives  quite  accurate  results  in  rapid 
blast-furnace  smelting,  and  where  abundance  of  iron  is  present.  If 
the  rate  of  smelting  be  slow,  and  considerable  lime  or  magnesia 
be  present,  5  per  cent,  of  the  sulphur  contents  of  the  charge  should 
be  deducted  before  beginning  the  calculation;  and  if  the  smelting 
furnace  is  a  reverberatory,  the  resulting  matte  will  average  8  per 
cent,  higher  in  copper  than  is  found  by  this  formula. 

A  simple  illustration  will  make  this  method  of  calculation  more 

We  will  assume  that  a  roasted  ore  having  the  following  composi- 
tion is  to  be  smelted  in  a  blast-furnace: 

*  This  calculation  refers  entirely  to  tlie  older  method  of  smelting,  without 
attempting  any  oxidizing  action  in  the  blast  furnace. 



Cu      =    9.0  per  cent.  Pb=  2.0  per  cent. 

Fe     =45.0        "  S*=7.8 

SiO,  •-=  27.0        "  O  and  loss  =7.2 

Zn     =    2.0        "  

Total,  100.00 



Following  the  rule  given, 

9  Cu  require  2  27  S  to  form  a  subsulphide. 
f  of  2  Pb  require  0.23  S  to  form  a  sulphide. 
^  of  2  Zn  require  0.50  S  to  form  a  sulphide. 

This  provides  for  3  per  cent,  of  the  7.8  per  cent,  of  sulphur 
present,  leaving  4.8  per  cent.,  which  will  take  up  enough  Fo  to 
form  a  raouosulphide.  Calculation  shows  that  8.4  per  cent,  of  Fe 
will  thus  be  required,  leaving  36.6  per  cent,  available  for  the  slag. 

In  order  to  express  the  composition  of  the  matte  just  calculated, 
in  the  ordinary  manner,  we  multiply  the  amount  of  each  ingredient 
by  a  common  factor  that  will  leduce  it  to  a  percentage.  In  this 
case  the  factor  is  3.61. 

9  Cu  +  2.27  S  =  11  27  X  3.61  =  40  09  per  cent.  Cu,S. 
1.5  Pb+ 0.23  S=  1.73X3.61=  6.25  "  PbS. 
l.Zn  +  0.50  S  =  1.5  X  3.61  =  5.41  "  ZnS. 
8.4  Fe+ 4.80  8  =  13.2    X  3.61  =  47.65        "         FeS. 

7.8  per  cent.  S  100.00 

Thus  the  matte  from  such  a  charge  will  contain  about  32.5  per 
cent,  copper;  the  slight  loss  of  sulphur  by  volatilization  and  as 
SO  being  usually  fully  balanced  by  the  presence  in  the  matte  of  a 
certain  proportion  of  subsulphides  in  place  of  sulphides,  or  even 
of  metallic  iron. 

The  same  charge  smelted  in  a  reverberatory  furnace  would  yield 
a  matte  of  about  40  per  cent.  Cu. 

The  proper  composition  of  tlie  slag  has  not  been  particularly 
considered  in  this  example.  It  would  be  somewhat  too  siliceous  for 
blast  furnace  work,  requiring  the  addition  of  a  little  limestone; 
while   for  reverberatory  work  it  would  be  about  right  as  it  stands. 

From  the  foregoing  statements  it  is  evident  that  in  ordinary 
copper  smelting  the  calcination   of  sulphide  ores  need  seldom  be 

*As  most  of  the  oxidized  compounds  of  sulphur  contained  in  the  calcined  ore 
will  be  reduced  to  sulphides  in  the  cupola  furnace,  it  is  proper  to  estimate  all 
the  sulphur  present  as  metallic  sulphur. 


piislied  to  the  point  of  perfectiou  indicated  when  treating  of  the 
chemical  reactions  that  take  place  in  the  roasting.  On  the  con- 
trary, a  due  regard  for  the  proper  quality  of  the  resulting  matte 
and  slag  will  probably  render  it  advisable  to  stop  the  calcining 
process  long  before  the  decomposition  of  the  sulphate  of  copper  in 
the  charge  is  complete,  and  even  while  a  considerable  portion  of 
nndecomposed  sulphides  still  remains.  If,  however,  the  calcina- 
tion has  been  carried  too  far,  it  is  very  easy  to  regulate  matters  by 
the  addition  to  the  smelting  mixture  of  a  small  proportion  of  raw 
snlphuret  ore. 

A  glance  at  the  behavior  of  the  various  compounds  of  sulphur 
and  bases  is  essential  for  the  clear  understanding  of  the  much 
greater  richness  of  the  matte  resulting  from  the  fusion  of  any 
given  charge  in  a  reverberatory  than  in  a  blast-furnace,  and  of  the 
importance  of  having  a  certain  proportion  of  sulphates  and  other 
oxidized  comiiounds  in  the  smelting  mixture,  in  order  that  they 
may  react  on  each  other  in  the  manner  best  calculated  to  eliminate 
the  residual  sulphur,  and  thus  in  a  measure  make  up  for  imiDerfect 

In  the  blast-furnace  but  little  sulphur  can  be  directly  volatil- 
ized, and,  consequently,  simply  fuses  with  the  copper  or  iron 
present  to  form  the  artificial  sulphide  called  matte.  But  the  sul- 
phates in  the  presence  of  carbonic  oxide  may  undergo  the  following 
reaction:  CO-fFeO,  SOs^COa+SO.+FeO;  the  carbonic  oxide 
burning  to  acid,  while  the  sulphuric  acid  is  reduced  to  sulphurous 
acid,  which  escapes  by  volatilization,  and  the  protoxide  ct  iron 
unites  with  silica  to  form  a  slag.  But  this  is  true  of  only  a  very 
small  proportion  of  the  sulphates  present,  as  in  the  powerful  re- 
ducing atmosphere  of  the  blast-furnace,  the  sulphurous  acid,  even 
when  once  formed,  comes  in  contact  with  an  overwhelming  pro- 
portion of  CO,  which  in  burning  to  CO2  robs  the  SO2  of  its  oxy- 
gen, reducing  it  to  sulphur,  in  which  condition  it  unites  with  iron 
or  copper  and  enters  the  matte,  thus  increasing  the  amount  of  this 
product,  while  it  robs  the  slag  of  its  most  valuable  constitueut. 
It  is  interesting  to  note  the  striking  ditlereuce  of  the  reaction  in 
the  reverberatory  furnace,  where  the  atmosphere  may  be  regarded 
as  neutral;  GO,  the  most  powerful  reducing  agent,  being  virtually 

CusS  +  4  CuO,  S05  =  6  CuO  -I-  5  SO2. 
Cu,S  +  2  CuO,  S03=  2  Cu,0  -f  3  S02. 
Cu=S  +  2  Cu,0         =  6  Cu      +  SO2. 


By  stiidyiug  these  formulaj,  it  will  uo  louger  seetn  strange  that 
the  revevberatory  produces  so  much  richer  matte  than  the  blast- 
furnace from  the  same  charge.  Nearly  all  the  reactions  between 
sulphides  and  sulphates  result  in  the  formation  of  oxides  and 
volatile  SO2,  and  were  it  not  for  an  almost  invariable  preponder- 
ance of  uudecomposed  sulphides  in  the  charge,  the  elimination  of 
the  sulphur  might  theoretically  be  almost  complete.  It  is  by  this 
all-important,  but  frequently  neglected,  establishment  of  a  proper 
proportion  between  the  sulphides  and  sulphates,  that  extraordi- 
nary results  may  be  obtained  in  reverberatory  smelting,  and  the 
roasting  plant  greatly  reduced,  as  shown  in  chapter  on  "Direct 
Method  of  Refining  Copper." 

Although  treating  of  smelting,  this  matter  belongs  strictly  to 
the  calcining  department,  and  presents  a  field  for  study  of  great 
interest  and  practical  value.  A  close  analogy  may  be  found  in  the 
various  reverberatory  processes  as  applied  to  the  smelting  of  galena 
ores,  where  almost  exactly  the  same  results  are  produced,  using 
lead  instead  of  copper,  and  obtaining  metallic  lead  with  a  minimum 
amount  of  calcination,  and  putting  to  accurate  practical  use  the 
reactions  just  explained,  although  text-books  on  copper  metallurgy 
are  strangely  silent  on  this  important  subject. 

The  length  of  tiuie  requisite  to  roast  a  charge  of  ore  of  a  given 
weight  in  the  long  furnace  under  discussion  depends,  of  course, 
upon  the  composition  of  the  charge  and  the  degree  of  thorough- 
ness in  oxidation  desired.  Each  of  the  four  iiearths  of  this  fur- 
nace has  an  effective  area  of  about  250  square  feet,  and  can  conse- 
quently receive  4,000  pounds  of  ore  if  only  16  pounds  to  the  square 
foot  are  charged.  This  is  a  very  moderate  charge,  especially  for 
heavy  sulphide  ores,  but  will  ordinarily  give  better  results  than  a 
heavier  burden.  It  will  cover  the  hearth  about  2^  inches  deep 
when  charged,  increasing  in  bulk  to  about  4  inches  at  the  comple- 
tion of  the  process.  By  shifting  each  charge  every  four  hours,  the 
ore  will  remain  16  hours  in  the  furnace,  a  time  generally  ample  to 
produce  the  desired  effect.  On  this  basis,  the  furnace  would  put 
through  12  tons  in  twenty-four  hours,  which  may  be  regarded  as 
its  maximum  capacity  on  su6h  ores  as  the  Butte  concentrates. 
But  this  is  the  extreme  limit  for  two  men  per' shift,  nor  will  these 
figures  be  reached  under  ordinary  circumstances.  Two  cords  of 
wood  or  2,240  pounds  of  soft  coal  should  supply  the  grate  for 
twenty-four  hours,  the  supply  of  air  to  the  ash-pit  being  kept  at 
the  lowest  possible  point.     The  sulphur  contents  of  the  ore  fur- 



nish  a  mnch  greater  proportion  of  the  beat  than  does  the  fuel  on 
the  grate. 

The  manipnlatious  pertaining  to  tlie  ordinary  calcination  of  ore 
are  too  simple  and  generally  known  to  be  worthy  of  a  place  in  a 
condensed  treatise. 

The  following  experiments  form  part  of  a  series  extending  over 
several  years.  The  author  desires  to  acknowledge  the  assistance 
of  Messrs.  J.  F.  Talbot  and  F.  Ames,  and  others,  in  the  chemical 
portion  of  the  work. 

Copper  in  Roasted  Ore. 

=  £ 

c  S 









*3    O 

c  *• 

r  t 












<  — 





Pr  ct.!  Lbs. 




37.0  '  4,130 








Heavy  pyritous  ore. 



39.0  1  4,130 






8.17  1 



Same  ore. 
\  Purple    ore    with 



31.0  1  3,925 





17.31  1 



•<     much  pyrites  and 
(     some  zincblende. 

4  .. 


31.0  1  3.940 








Same  ore. 



24.3  :  3,600 







Matte  from  cupola. 



22.0  '  3,580 





64.30  1 


( Blue     metal    from 
1     reverberatorv. 

7. . . 


21.4     3,800 





74.90  ! 


\  White   metal  from 
\     reverberatory. 

The  loss  of  weight  from  the  removal  of  the  sulphur  is  partially 
balanced  by  the  oxygen  combining  with  the  metallic  bases,  and  is 
exceedingly  variable,  as  may  be  seen  by  this  table. 

The  loss  in  copper  during  calcination  is  very  small,  and  almost 
entirely  mechanical,  being  for  the  most  part  recoverable  where 
proper  arrangements  are  made  for  the  deposition  of  the  flue-dust. 
Average  results  from  personal  experience  show  a  loss  of  about  1:^ 
per  cent,  of  the  original  copper  contents  of  the  ore  during 

This  flue-dust  is  usually  of  very  much  lower  grade  than  the  ore 
from  which  it  results,  being  diluted  with  the  dust  from  the  fluxes, 
fuel,  etc.,  and  generally  contains  from  20  to  30  per  cent,  of  its 
value  in  a  soluble  form,  thus  prohibiting  the  use  of  water  as  an 
aid  to  its  condensation,  unless  provision  is  made  to  precipitate  the 
dissolved  metal. 

Unless  the  ores  treated  are  of  remarkable  purity,  it  is  best  to 
smelt  the  flue-dust  by  itself.  Otherwise,  the  quality  of  the  metal 
is  likely  to  suffer,  as  the  substances  most  injurious  to  it — arsenic, 
antimony,  and  tellurium — are  volatile  and  sure  to  be  condensed  in 
the  flues,  thus  being  collected  in  a  concentrated  form. 



The  various  methods  employed  iu  roasting  (calciuirig)  copper 
ores  fall  naturally  into  two  principal  divisions,  according  to  the 
mechanical  condition  of  the  ore,  whether  fine  or  coarse. 

The  main  divisions  may  be  again  subdivided,  according  to  the 
means  employed  for  executing  the  operation  of  roasting.  The 
following  Scheme  makes  a  convenient  working  classification: 

{A)  Roasting  ores  in  lump  form. 

1.  Heap  roasting.     Suited  to  both  ore  and  matte. 

2.  Stall  roasting: 

(«)  Open  stalls.     Suited  only  to  ore. 

{b)  Covered  stalls.     Suited  to  both  ore  and  matte. 

3.  Kiln  roasting.     Suited  only  to  ore. 
(B)  Roasting  ores  in  pulverized  condition. 

1.  Shaft  furnaces. 

2.  Stalls. 

3.  Hand  reverberatory  calciners. 

(a)  Open  hearth. 
{b)  Muffle. 

4.  Revolving  cylinders. 

(a)  Continuous  discharge. 

[b)  Intermittent  discharge. 

5.  Automatic  reverberatory  calciners. 

(a)  Stationary  hearth. 
{b)  Revolving  hearth. 

Coarse  ore  that  comes  from  the  mine  in  pieces  of  varying  vol- 
ume must  be  broken  to  a  proper  maximum  size  for  the  operation 
that  it  is  to  undergo.  This  size  varies  so  greatly  with  local  con- 
ditions that  it  is  impossible  to  lay  down  any  exact  rules  on  the 
subject.  The  matter  will  be  considered  separately  for  each 

As  we  invariably  have  more  fines  than  we  require,  or  can  use, 
for  the  covering  of  the  heap  in  ore,  or  stall-roasting,  it  follows 


that  economy  warus  us  to  go  no  further  in  the  crushing  than  is  abso- 
lutely essential  for  the  success  of  the  roasting.  By  crushing  an 
ore  any  finer  than  this  we  lose  in  four  different  ways: 

1.  In  the  extra  cost  of  fine  crushing. 

2.  In  dust. 

3.  In  hampering  the  subsequent  smelting  process.     (This  ap- 

plies only  where  the  ore  is  to  be  smelted  in  blast  furnaces.) 

4.  In  tlie  increased  expense  incurred  in  roasting  the  excess  of 

fines  in  appropriate  furnaces,  or  of  using  them  half  raw  in 
the  smelting  furnaces. 

The  second  and  fourth  of  these  objections  have  been  practically 
canceled  by  the  introduction  of  automatic  calciners  that  operate  so 
quietly  that  the  loss  from  dust  is  less  than  in  heap  roasting,  while 
the  cost  per  ton  of  the  operation  will  not  exceed  that  of  the  ruder 
method.  Even  where  wages  are  very  low  and  heap  roasting  is 
cheap,  the  thoroughness  and  uniformity  of  the  furnace  operation 
will  often  render  the  latter  more  economical.  But,  as  the  heap  or 
stall  yields  a  coarse  product  admirably  suited  to  blast-furnace  work, 
and  also  avoids  the  heavy  outlay  for  a  calcining  plant,  it  will,  no 
doubt,  long  be  used  in  remote  districts,  and  in  the  early  stages  of 
certain  metallurgical  enterprises. 

Heuce,  it  is  of  importance  to  crush  the  ore  intended  for  heap, 
stall,  or  kiln-roasting,  in  such  a  manner  as  to  make  the  smallest 
possible  proportion  of  fines,  providing,  always,  that  the  method 
pursued  is  sufficiently  economical  and  rapid. 

Ores  containing  a  high  percentage  of  sulphur — 25  and  over 
— will  give  excellent  results  if  so  broken  that  the  largest  frag- 
ments shall  be  capable  of  passing  through  a  ring  3  inches  in 
diameter,  and,  in  some  cases,  will  roast  to  perfection,  if  sufficient 
time  be  given,  in  lumps  the  size  of  a  man's  head,  while  more  rocky 
ore,  which  is  likely  to  be  of  a  harder  and  denser  texture,  should 
be  reduced  to  pass  a  2-inch  opening.  Careful  experiments  can 
alone  determine  the  most  profitable  size  for  any  given  material, 
and  should  be  continued  on  a  large  scale  until  the  metallurgist  in 
charge  has  fully  satisfied  himself  on  this  point.  This  may  be  de- 
termined with  the  least  trouble  and  expense  by  noticing  the  weight 
and  quality  of  the  matte  obtained  by  smelting  the  roasted  ore  from 
various  heaps  formed  of  fragments  diifering  in  their  maximum 

All  other  conditions  being  identical,  the  heap  that  yields  the 
smallest  quantity  of  the  richest  matte  has,  of  course,  undergone 


the  most  perfect  oxidatiou,  aurl  should  be  selected  as  a  standard 
for  future  operations.  Variations  that  may  occur  in  the  chemical 
or  mechanical  condition  of  the  ore  should  be  carefully  watched  as 
a  guide  in  fixing  upon  the  best  roasting  size.  Local  conditions 
must  determine  whether  a  jaw-crusher  or  hand  labor  should  be 
used  for  this  purpose.  The  production  of  fines  is  a  decided  evil  in 
the  preparation  of  ore  for  heap  roasting,  and  the  manual  method 
possesses  a  certain  advantage  in  this  respect,  thougli  this  consider- 
ation may  be  outweighed  by  other  economic  conditions.  A  trial 
of  the  comparative  amount  of  fines  produced  by  machine  and 
liand-breaking  was  carried  out  on  three  diiferent  varieties  of  sul- 
phureted  copper  ores  of  average  hardness,  and  aggregating  2,220 
tons.*  The  broken  ore  was  thoroughly  screened;  all  passing 
through  a  sieve  of  three  meshes  to  the  inch  (0  mm.  openings)  was 
-designated  as  fines. f  One-half  (1,110  tons)  of  this  material  was 
passed  through  a  seven  by  ten  jaw  rock-breaker,  with  corrugated 
crnshing-plates  (which  produce  a  decidedly  less  projjortion  of  fines 
than  the  smooth  plates).  The  breaker  made  .240  revolutions  per 
minute,  and  had  a  discharge  opening  of  2|  inches.  The  other 
moiety  was  broken  by  experienced  Avorkmon,  with  proper  spalling- 
hammers,  into  fragments  of  a  similar  maximum  size.  The  result 
was  as  follows,  only  the  fine  product  being  weighed,  the  coarse 
being  estimated  by  subtracting  the  former  from  the  total  amount: 


Tons.  Per  cent. 

Fine  product — below  6  mm.  in  diameter 192.25  17.32 

Coarse  product — between  6  mm.  and  64  mm. . . .  917.75  82.68 

Total 1,110.00        100.00 


Fine  product — below  6  mm.  in  diameter 103.34  9.31 

Coarse  product — between  6  mm.  and  64  mm.  ..1,006.66  90.69 

Total 1.110.00         100.00 

*  Unless  otherwise  indicated,  all  tons  equal  2,000  pounds. 

f  It  should  also  be  explained  that,  owing  to  the  large  and  very  variable 
amount  of  tine  material  contained  in  the  ore  before  crushing,  as  it  came  from 
the  mine,  it  was  passed  over  the  screen  just  referred  to  before  being  either  fed 
to  the  crusher  or  spalled  by  hand.  Without  this  precaution,  the  results  of  the 
trial  would  have  been  valueless,  as  the  variation  in  the  amount  of  fines  in  the 
original  ore  was  far  greater  than  the  discrepancy  in  the  amount  produced  bv 
the  two  different  methods  of  crushing. 


These  results  are  quite  in  accordance  with  impressions  derived 
from  general  observation,  and,  as  will  be  noticed,  prove  that,  with 
certain  classes  of  ore,  mechanical  crushing  produces  nearly  double  as 
much  fines  as  hand-spalliug.  As  10  per  cent,  of  fines  is  an  ample 
allowance  to  form  a  covering  for  any  kind  of  roast-heap — and 
better  results  are  obtained  when  the  same  partially  oxidized  mate- 
rial is  used  over  and  over  again  as  a  surface  protection — it  may 
frequently  occur  that,  in  spite  of  its  greater  cost,  haud-spalling 
may  prove  more  profitable  than  machine-breaking.  This  is  a 
matter  for  individual  decision,  and  can  be  determined  only  after  a 
mature  consideration  of  the  difference  in  expense  of  the  two  opera- 
tions, the  means  at  hand  for  the  calcination  and  subsequent 
advantageous  smelting,  of  the  increased  quantity  of  fines,  and 
whatever  other  factors  may  bear  on  the  case  in  hand.  The  follow- 
ing steps  should  be  carried  out,  whichever  method  is  decided  upon. 
The  ore,  after  breaking,  should  be  separated  into  three  classes, 
the  largest  including  all  the  product  between  the  maximum  size 
and  one  inch  (25  mm.);  the  medium  size,  or  ragging,  consisting 
of  the  class  between  25  mm.  and  the  fine  size  (three  meshes  to  the 
inch,  which  would  give  openings  of  about  6  mm.  net);  and  the 
fines,  as  already  explained.  Eoughly  speaking  the  percentage  of 
each  class,  including  the  fine  ore  that  is  invariably  produced 
during  the  operation  of  mining,  may  be  represented  by  the  fol- 
lowing figures: 

Coarse 55  per  cent. 

Ragging . .    ....    25         " 

Fines '. 20 

Total 100 

This  classication  is  effected  with  great  ease  and  economy  in  case 
machine-breaking  is  decided  upon,  by  the  use  of  a  cylindrical  or 
conical  screen  of  ^-inch  boiler  iron,  about  10  feet  in  length  and 
48  inches  or  more  in  diameter.  This  is  plsiced  below  the  breaker 
so  that  it  receives  all  the  ore.  It  is  made  to  revolve  from  12  to  10 
times  per  minute,  and  has  a  maximum  fall  of  an  inch  to  the  foot. 
This  can  easily  separate  20  tons  of  ore  per  hour,  and  by  proper 
arrangement  of  tracks  or  bins,  discharge  each  class  into  its  own 
]iin.  One  fault  in  this  very  simple  classifying  apparatus  is, 
that  the  coarse  lumps  of  ore  must  necessarily  traverse  all  the- 
finer  sizes  of  screen,  thus  greatly  augmenting  the  wear  and   tear- 


This  objection,  though  frequently  valid  iiDcler  other  circumstances, 
has  but  little  weight  when  it  is  remembered  that  even  the  smallest 
holes  (6  mm.)  are  punched  in  iron  of  such  thickness  {^  inch)  that 
it  will  withstand  even  the  roughest  usage  for  many  mouths.  To 
produce  the  three  sizes  just  alluded  to,  the  screen  requires  two 
sections,  with  holes  respectively  6  mm.  and  25  mm.,  of  which  the 
finer  size  should  occupy  the  upper  6  feet,  and  the  coarser  the 
lower  4  feet  of  the  screen.  In  remote  districts,  where  freight  is 
one  of  the  principal  items  of  expense,  heavy  iron  wire  cloth  may 
be  substituted  for  the  punched  boiler  iron,  and  if  properly  con- 
structed and  of  sufficiently  heavy  stock,  will  be  found  satisfactory, 
lasting  about  one-half  as  long  as  the  more  solid  material.  The 
difference  in  size  between  a  circular  hole  25  mm.  in  diameter  and  a 
square  with  sides  of  that  length,  should  not  be  overlooked  in 
changing  from  one  variety  of  screen  to  the  other.  The  mouth  of 
the  crusher  should  be  level  with  the  feeding-floor,  and  the  latter 
should  be  covered  with  quarter-inch  boiler  iron,  firmly  attached 
to  the  planks  by  countersunk  screws,  by  which  arrangement  the 
shoveling  is  greatly  facilitated.  With  such  a  plant,  three  good 
laborers  will  feed  the  breaker  at  the  rate  of  20  tons  an  hour  for  a 
10-hour  shift,  provided  none  of  the  rock  is  in  such  masses  as  to 
require  sledging,  and  that  the  ore  is  dumped  close  to  the  mouth  of 
the  breaker.  A  nine  by  fifteen  jaw-breaker  of  the  best  and  heaviest 
make  is  capable  of  crushing  the  amount  just  mentioned  to  a  maxi- 
mum size  of  2|  inches,  provided  the  rock  is  brittle,  heavy,  and 
not  inclined  to  clog  the  machine. 

The  expense  per  ton  of  breaking,  sizing,  and  delivering  into 
cars  with  such  a  plant,  operating  upon  ores  of  medium  tenacity,  is 
as  follows,  the  figures  being  deduced  from  average  results  of  han- 
dling large  quantities  under  the  most  varying  conditions.  It  is 
assumed  that  the  breaker  is  run  by  an  independent  engine  of  suffi- 
cient power,  while  the  wages  of  an  engineer  and  firemen  are  par- 
tially saved  by  taking  the  steam  from  the  boilers  that  are  supposed 
to  supply  the  main  works: 


COST     OF     BREAKING     OKE     BY     MACHINERY     WITH    A  PLANT    OF     200     TONS 

Per  shift.       Per  ton. 
Power — per  day  of  10  hours,  at  one  cent  per  ton  . .  .$2.00        $0.0100 

Labor : 
Four  laborers  at  $3.00 12.00  0.0600 

Repairs : 

Toggles  and  jaw  plates,  etc $0.85 

Wear   of    tools.    Babbitt    for    renewing 

bearings,  etc 0.7.5 

Daily  slight  repairs  on  machinery 0.80 

Miscellaneous  items,  sampling  etc 0.75  3.15  0.0157 

Sinking  Fund : 

To  replace  machinery  at  10  per  cent,  on 
original  cost 1.40  0.0070 

Total $18.55        $0.0928 

If  it  should  seem  at  first  glance  that  10  cents  per  ton  is  an  au- 
reasonably  low  figure,  it  will  be  noticed  that  the  cost  of  transpor- 
tation both  to  and  from  the  breaker  is  not  included  in  this  estimate; 
the  former  is  usually  charged  to  mining  expenses,  and  the  latter 
to  heap-roasting.  Ore  that  is  to  undergo  roasting  in  kilns  for  the 
purpose  of  acid  manufacture  must  be  broken  considerably  smaller 
than  that  just  described,  and  this,  of  course,  lessens  the  capacity 
of  the  apparatus  and  proportionately  increases  the  expense.  An 
increase  of  50  per  cent,  in  the  above  estimate  will  be  sufficient  to 
cover  it.  The  figures  given  above  have  been  frequently  attained 
bv  the  author,  but  only  under  certain  favorable  conditions,  among 
which  are:  Abundance  of  power  to  run  the  breaker  to  its  full 
speed,  regardless  of  forced  feeding.  A  constant  sytem  of  supervi- 
sion by  which  the  plant  is  kept  up  to  its  full  capacity  of  20  tons 
per  hour,  and  which  demands  exceptionally  good  men  as  feeders. 
A  frequent  inspection  of  the  machinery,  and  renewal  of  all  jaw 
plates,  toggles,  and  other  wearing  parts,  before  the  efficiency  of 
the  machine  has  begun  to  be  impaired;  all  of  which  repairs  should 
be  foreseen  and  executed  during  the  night  shift  or  on  idle  days. 
A  perfect  system  of  checking  the  weight  of  all  ore  received  and 
crushed,  without  which  precaution  a  mysterious  and  surprisingly 
large  deficit'will  be  found  to  exist  on  taking  stock.  It  is  hardly 
necessary  to  mention  that  all  bearings  that  cannot  be  reached  while 
the  machinery  is  in  motion  must  be  provided  with  ample  self- 
oilers,  and   since  clouds  of  dust  are  generated  in   this  work,  that 


unusual  care  must  be  taken  in  covering  and  protecting  all  boxes 
and  parts  subject  to  injury  from  this  cause.  Unless  the  ore  is 
sufficiently  damp — either  naturally  or  by  artificial  sprinkling — to 
prevent  this  excessive  production  of  dust,  the  feeders  should  be 
required  to  wear  some  efficient  form  of  respirator;  otherwise,  they 
are  likely  to  receive  serious  and  permanent  injury,  the  fine  parti- 
cles of  sulphides  being  peculiarly  irritating  to  the  lungs  and  entire 
bronchial  mucous  membrane. 

The  ireaking  of  ore  by  hand  hammers^  technically  denominated 
"spalling,"  is  worthy  of  more  careful  consideration  than  is  gener- 
jilly  bestowed  upon  it.  The  style  of  hammer  is  seldom  suited  to 
the  purpose,  though  both  the  amount  of  labor  accomplished  and 
the  personal  comfort  of  the  workmen  depend  more  upon  the  weight 
and  shape  of  this  implement  and  its  liandle  than  on  any  other 
single  factor  save  the  quality  of  the  ore  itself.  There  should  be 
several  cast-steel  sledges,  dilfering  in  weight  from  6  to  14  pounds, 
and  intended  for  general  use  in  breaking  up  the  larger  fragments 
of  rock  to  a  size  suitable  for  the  light  spalling-hammers.  Each 
laborer  should  be  provided  with  a  hammer  6  inches  in  length, 
forged  from  a  1^-inch  octagonal  bar  of  the  best  steel,  and  weighing 
about  2f  pounds.  This  should  be  somewhat  flattened  and  expanded 
at  the  middle  third,  to  give  ample  room  for  a  handle  of  sufficient 
size  to  prevent  frequent  breakage.  The  handles  usually  sold  for 
this  purpose  are  a  constant  source  of  annoyance  and  expense,  being 
totally  unsuited  to  this  peculiar  duty.  It  is  better  to  have  the 
handles  made  at  the  works,  if  it  is  possible  to  procure  the  proper 
variety  of  oak,  ash,  hickory,  or,  far  better  than  all,  a  small  tree 
known  in  New  England  as  ironwood  or  hornbeam,  which,  when 
peeled  and  used  in  its  green  state,  excels  most  other  woods  in 
toughness  and  elasticity.  The  handles  should  be  perfectly  straight, 
without  crook  or  twist,  so  that,  when  firmly  fastened  in  the  eye  of 
the  hammer  by  an  iron  wedge,  the  hammer  hangs  exactly  true. 
Their  value  and  durability  depend  much  upon  the  skill  with  which 
the  handles  are  shaved  down  to  an  area  less  than  half  their  maxi- 
mum size,  beginning  at  a  point  some  six  inches  above  the  hammer- 
head and  extending  for  about  ten  inches  toward  the  free  extremity. 
If  properly  made  and  of  good  material,  they  may  be  made  so  small 
as  to  appear  liable  to  break  at  the  first  blow;  but  in  reality  they 
are  so  elastic  that  they  act  as  a  spring,  and  obviate  all  disagreeable 
effects  of  shock,  wear  longer, and  do  more  work  than  the  ordinary 
handle.     Such  a  handle  has  lasted  five  months  of  constant  use,  in. 


the  hands  of  a  careful  ■workman,  whereas  one  of  the  ordinary  make 
has  an  average  life  of  scarcely  four  days,  or  perhaps  thirty  tons  of 
ore.  Where  the  ore  is  of  pretty  uniform  character,  it  is  advan- 
tageous to  adopt  the  contract  system  for  this  kind  of  work.  A 
skillful  laborer.,  under  ordinary  conditions,  will  break  seven  tons 
of  rock  per  10-hour  shift  to  a  size  of  2^  inches,*  taking  coarse 
and  fine  as  it  comes,  and  in  some  cases  he  is  also  able  to  assist  in 
screening  and  loading  the  same  into  cars.  This  latter  operation 
should  be  executed  with  a  stroug  potato-fork  having  such  spaces 
between  the  tines  as  to  retain  the  coarsest  size,  while  the  finer 
classes  are  left  upon  the  ground.  These  forks  are  made  for  this 
purpose  by  a  firm  in  St.  Louis,  and  are  much  superior  to  the  ordi- 
nary forks.  When  a  sufficient  quantity  of  the  finer  classes  lias 
accumulated  and  the  pile  or  stall  is  ready  to  receive  its  outer  layer 
of  ragging,  the  mixed  material  should  be  thrown  upon  a  screen 
inclined  to  an  angle  of  about  48  degrees  and  having  three  meshes 
to  the  inch.  This  screen  is  elevated  upon  legs  to  such  a  height 
that  the  coarser  class  that  fails  to  pass  its  openings  will  be  caught 
in  a  car  or  barrow,  while  the  fines  fall  either  into  a  second  movable 
receptacle  or  upon  the  floor,  being  in  the  latter  case  prevented 
from  again  mixing  with  the  unscreened  ore  by  a  tight  boarding  on 
the  front  and  sides  of  the  screen  frame.  The  amount  of  space 
required  for  convenient  spalling  is  about  40  square  feet  per  man, 
which  will  allow  for  ore-dumps,  tracks,  sample  boxes,  etc.  A 
good  light  is  essential,  especially  if  any  sorting  is  to  be  done,  and 
it  is  in  this  case  and  where  fuel  is  expensive  that  haud-spalling 
frequently  presents  especial  advantages.  When  the  ores  are  sili- 
ceous, a  mere  rejection  of  such  pieces  of  barren  quartz  or  wall  rock 
as  have  accidentally  got  among  the  ore,  or  first  become  visible  on 
breaking  up  the  larger  masses,  may  have  a  most  beneficent  influ- 
ence on  the  subsequent  fusion.  Where  the  expense  of  treatment 
is  high  and  work  is  conducted  on  a  large  scale,  the  profit  resulting 
from  raising  the  average  contents  of  the  ore  even  a  single  per  cent, 
is  hardly  credible,  even  aside  from  the  increased  fusibility  due  to 
the  diminished  proportion  of  silica. 

The  cost  of  spalling  an  ore  of  the  same  character  as  that  on 
which  the  foregoing  estimates  for  machine-breaking  are  based,  has 
been  calculated  from   the  average  results  of  a  very  large  quantity 

*  Unless  otherwise  specified,  the  term  "  day  "  or  "  shift  "  may  be  understood 
to  signify  the  ordinary  working  day  of  ten  hours,  from  seven  a.m.  to  six  p.m., 
with  one  hour  for  dinner. 


of  ore,  assuming  100  tons  to  be  spalled,  screened,  and  loaded  in 
ten  hours. 

COST     OP     SPALLING     ORE     BY     HAND     WITH    AN     OUTPUT    OF   100  TONS   PER 


Labor :  * 

Per  100  tons.  Per  tou. 
L4  men  breaking  ore,  including  screen- 
ing and  loading,  at  |1.50 |21.00 

4  men  sledging  and  loading  at  $1.50. .. .     6.00 

1  foreman 3,50        $29.50        $0,295 

Repairs  : 

Including  new  steel  and  handles. 

5  handles  at  30  cents 1.50 

7  pounds  of  steel  at  15  cents 1.05 

Blacksmith's  and  other  work  on  above, 

1^  day 1.00 

Screens,  forks,  and  shovels 1.67 

General  repairs  0.55  5.77        0.0577 

Sinking  Fund  : 

To  replace  screens  and  permanent  fix- 
tures     0.15        0.0015 

Total $35.42     $0.3542 

Perhaps  the  most  marked  point  of  difference  between  the  roast- 
ing of  lumps  and  fines  is  the  time  requisite  for  their  oxidation. 
Oxidation  is  almost  instantaneous  for  an  infinitely  small  particle 
of  any  sulphide,  and  the  time  increases  with  the  cubic  contents  of 
the  fragment,  until  such  a  size  is  reached  that  the  air  fails  to 
penetrate  the  thick  crust  of  oxides  formed  upon  the  outside  of  the 
lump  of  ore  or  matte,  and  all  action  ceases. 

It  might  seem,  therefore,  that  the  process  of  pulverization 
should  be  pushed  to  extreme  limits,  and  that  the  best  results 
would  be  obtained  from  the  most  finely  ground  ore.  But  this  is 
by  no  means  the  case  in  actual  practice;  for  other  conditions  arise 
that  more  than  counteract  any  advantage  in  time.  The  chief  of 
these,  aside  from  the  difficulty  and  expense  involved  in  the  pro- 
duction of  such  fine  pulp,  are  the  losses  in  metal,  both  mechanical 
and  chemical,  that  occur  with  every  movement  of  the  ore,  and 
reach  an  enormous  aggregate  before  the  operation  is  completed; 
and  the  liability  to  fritting  or  sticking  together  in  the  calcining- 
furnace,  regardless  of  the  greatest  possible  care  in  this  process. 

■*  I  assume  that  wages  are  low;  otherwise   machine  breakers  would  be  used. 


The  o.xidation  of  the  particles  takes  place  with  such  rapidity  that 
a  temperature  is  generated  above  the  fusion-point  of  ordinary 

Still  further  objections  could  be  mentioned;  but  those  already 
adduced  are  sufficient  to  limit  the  degree  of  pulverization  for  the 
principal  portion  of  the  ore,  although  a  greater  or  less  proportion, 
according  to  the  machinery  used  for  the  purpose,  is  crushed  to  an 
impalpable  dust,  and  causes  a  considerable  mechanical  loss,  in 
spite  of  all  provision  for  its  prevention. 

The  best  size  to  which  to  crush  varies  with  each  individual  ore, 
and  is  entirely  a  matter  of  trial  and  experience;  nor  should  any 
one  responsible  for  the  calcination  of  any  given  material  rest  satis- 
fied until  he  has  determined  by  actual  and  long-continued  experi- 
ment, that  the  substitution  of  either  a  coarser  or  a  finer  screen  for 
the  size  in  use  will  be  followed  by  less  favorable  results. 

This  may  be  arrived  at  by  careful  comparative  determinations 
of  the  residual  sulphur  contents  after  the  calcination  of  material 
crushed  through  screens  of  various  sized  mesh  and  roasted  for  the 
same  length  of  time,  careful  consideration  also  being  given  to  the 
cost  of  crnshiug  in  each  case,  to  the  condition  of  the  oxides  of  iron 
present  (the  sesquioxide  is  an  unfavorable  constitutent  in  rever- 
beratory  smelting),  and,  above  all,  to  the  quantity  of  flue-dust 
formed,  and  loss  of  metal  by  volatilization. 

It  is  evident  that  such  diverse  and  obscure  questions  can  only 
be  accurately  determined  by  extensive  and  long-continued  trials. 
But  the  result  is  well  worth  the  labor,  and  in  these  days  of  almost 
universal  information  and  close  competition,  it  is  only  by  such 
means  that  any  decided  advantage  can  be  obtained. 

While  mattes,  speiss,  or  similar  products  of  fusion  must  always 
be  granulated  or  pulverized  to  the  degree  required  for  calcination, 
it  is  not  an  uncommon  quality  of  sulphide  ores  either  to  decrepi- 
tate, or  else  to  fall  to  pieces  when  heated  by  the  mere  moving 
from  place  to  place  in  ^.he  furnace,  to  such  an  extent  that  the 
charge  may  be  made  up  f  pieces  from  the  size  of  a  walnut  down, 
without  affecting  either  the  time  requisite  for  the  oxidation  or  for 
its  perfection.  The  product  will  be  an  almost  homogeneous  and 
impalpable  powder. 

A  more  striking  illustration  of  such  a  condition  of  affairs  can 
hardly  be  found  than  in  the  case  of  the  concentrates  from  the 
Parrot    Company's  mine  at  Butte,  Montana. 

In  this  instance,  the  process  of  subdivision  resulted  from   two 


different  causes.  The  iron  pyrites  that  forms  the  larger  portion  of 
the  ore  decrepitates  into  very  minute  cubes,  which  are  subsequently 
reduced  to  a  fine  powder  by  oxidation,  while  the  fragments  of  pure 
copper  ore — ^bornite — seem  gradually  to  diminish  in  size  by  the 
wearing  away  of  the  surface  as  it  becomes  earthy  and  friable  from 
the  superficial  formation  of  oxides. 

This  latter  phenomenon  may  also  be  observed  to  a  less  extent  in 
the  calcination  of  mattes  when  they  are  of  a  sufficiently  soft  or 
porous  nature;  but  in  roasting  a  considerable  quantity  of  a  very 
low-grade  matte  (from  10  to  15  per  cent,  of  copper)  that  had  been 
©btained  in  hard  polished  granules  by  tapping  into  water,  it  was 
found  impossible  materially  to  alter  either  the  size  or  shape  of  the 
grains,  many  of  which  were  as  large  as  an  army  beau,  or  satisfac- 
torily to  reduce  the  percentage  of  sulphur,  even  by  long  exposure 
to  a  temperature  closely  approaching  its  fusion-point. 

On  the  other  hand,  quite  satisfactory  results  are  obtained  in  the 
case  of  richer  matte  (from  30  to  40  per  cent,  of  copper)  by  granu- 
lation in  water;  and,  in  many  of  the  foreign  works,  this  is  the 
only  means  provided  for  the  preparation  of  the  matte  for  the  pro- 
cess of  roasting;  but  it  must  be  remembered  that  this  practice  is 
confined  to  the  English  reverberatory  method,  where  it  is  not 
desired  to  remove  more  than  50  per  cent,  of  the  sulphur  by  roast- 
ing, and  where  a  portion  of  sulphides  still  remains  in  the  calcined 
matte  that  would  be  entirely  nnsuited  to  the  so-called  "blast- 
furnace" method  of  matte  concentration  in  cupolas,  as  usually 
practiced  in  this  country. 

Although  the  results  described,  as  obtained  by  granulation,  may 
be  improved  upon  by  careful  attention  to  the  temperature  and 
pitch  of  the  matte  when  tapped,  and  especially  by  care  and  experi- 
ence on  the  part  of  the  smelter,  this  practice  cannot  be  recom- 
mended, excepting  under  peculiar  conditions  and  in  remote  situa- 
tions where  improved  crushing  machinery  is  not  obtainable,  or 
where  the  physical  condition  of  the  matte  is  particularly  favorable 
to  the  production  of  porous  and  friable  granules.  Nor  is  anvthing 
gained  by  its  employment  for  the  purpose  of  avoiding  the  prepara- 
tory breaker,  and  obtaining  at  once  a  material  sufficiently  subdi- 
vided for  immediate  treatment  in  the  final  pulverizing  apparatus; 
for,  although  in  this  practice  the  larger  granules  are  broken  and 
crushed  into  a  condition  favorable  for  the  calcining  process,  a  large 
proportion  of  the  entire  mass  is  already  so  small  as  to  pass 
through  tiie  crushing  apparatus  untouched,  in  the  shape  of  minute 


spherical  pellets  or  globules  which  present  the  least  possible  surface 
to  oxidation,  and  retain  a  hard,  glossy  surface.  These  grains  art 
scarcely  aSected  by  any  moderate  temperature,  and  may  even 
undergo  complete  fusion  without  any  perceptible  loss  of  sulphur. 
Not  many  years  ago,  jhe  question  of  economy  might  have  influ- 
enced the  adoption  of  this  practice;  but  at  the  present  time,  and 
in  view  of  the  improved  and  comparatively  inexpensive  machinery 
at  our  disposal,  it  is  probable  that  the  inconvenience,  danger,  and 
other  drawbacks  inseparable  from  the  projection  of  large  quanti- 
ties of  molten  sulphides  into  water,  and  their  subsequent  recovery 
from  the  reservoir  or  whatever  vessel  is  employed  for  the  purpose, 
more  than  outweigh  the  cost  of  crushing  by  machinery. 

It  is  impossible  to  lay  down  fixed  rules  for  the  degree  of  pulver- 
ization of  any  material  best  suited  to  roasting.  Each  case  must 
be  decided  according  to  its  own  peculiar  conditions,  including  the 
cost  of  labor  and  power,  and  the  capacity  and  quality  of  the 
mechanical  means  available. 

Bearing  in  mind  the  results  that  may,  in  certain  exceptional 
cases,  come  from  decrepitation,  it  may  be  assumed  that  reduction 
in  size  beyond  one-eighth  of  an  inch  is  seldom  advantageous  in 
treating  ores,  and  that  the  presence  of  a  large  proportion  of  sul- 
phides, or  of  a  particularly  porous  or  friable  gangue,  may  permit 
an  increase  of  the  screen  mesh  to  one-fourth  inch  or  more.  With 
mattes,  a  slightly  finer  standard  (from  one-sixth  to  one-eighth 
inch)  may  be  employed. 

The  proportion  of  the  ore  reduced  to  a  minuteness  neither  in- 
tended nor  desired,  depends  materially  upon  the  means  employed 
for  crushing;  and  as  the  mechanical  loss  and  other  evils  enumer- 
ated increase  in  direct  ratio  to  the  amount  of  fine  dust  in  the 
charge,  it  is  evident  that,  other  things  being  equal,  the  apparatus 
best  adapted  to  the  breaking  of  ore  or  matte  is  that  which  produces 
the  smallest  proportion  of  fines. 


The  crushing  machinery  used  for  the  purpose  under  discussion 
may  be  divided  into  two  classes. 

1.  For  preparatory  crushing:  Breakers  of  various  patterns. 

2.  For  final  pulverization:  Stamps,  Ball  pulverizers,  Chili  mills, 
various  patent  pulverizers  and  grinders,  Cornish  rolls. 



Apart  from  machines  intended  for  fracturing  large  masses  of 
rock,  such  as  rock-hammers,  all  preparatory  breakers  consist  essen- 
tially of  a  movable  Jaw  that  squeezes  the  pieces  of  rock  against  a 
more  or  less  rigid  frame.    The  most  useful  types  are: 

The  Gates  crusher,  and 

'I  he  Comet  crusher,  in  both  of  which  the  movable,  jaw  is  a 
massive  cone,  suspended,  or  supported  on  a  step,  which  oscillates 
with  a  gyrating  motion  in  a  heavy,  bottomless,  cup-shaped  mortar. 
This  type  has  very  great  capacity. 

The  universally-known  Blake  crusher,  of  which  there  are  several 
patterns,  has  a  movable  jaw  hanging  from  two  pivots,  which  is 
pressed  against  the  stationary  one  by  a  pitman,  or  by  a  vertical 
connecting-rod  and  toggles. 

The  Blake-Challenge  is  a  sectional  machine,  constructed  of 
wrought  iron  and  steel,  the  heavy  thrust  of  the  crushing  being 
taken  by  two  powerful  steel  rods. 

The  Blake  multiple-jaw  breaker  is  considered  in  the  following 

The  Dodge  crusher  is  particularly  suited  to  reducing  fragments 
of  moderate  size  to  still  finer  grains,  and  when  built  of  sufficient 
strength  to  admit  of  the  wide  jaw  that  is  necessary  for  large  capac- 
ity, is  a  most  useful  machine.  The  jaw  oscillates  on  fixed  points 
projecting  from  its  lower  extremity,  the  maximum  of  motion 
being  at  the  top  of  the  jaw. 

There  are  other  excellent  patterns  of  jaw-crnshers  that  are 
suited  to  special  conditions. 

The  principle  of  jaw-crushing  is  eminently  satisfactory  as  re- 
gards economy,  capacity,  and  general  suitability  to  the  purpose  for 
which  it  is  intended. 

A  machine  should  be  selected  that  has  stood  the  test  of  years, 
and  is  manufactured  by  some  well-known  and  reputable  firm. 
Light-built  machines  should  be  particularly  avoided,  as  the  strain 
exerted  upon  certain  parts  of  every  breaker,  especially  when  clogged 
with  clayey  ore  and  set  to  crush  fine  without  shorteni«g  the  stroke 
of  the  jaw,  is  something  enormous,  and  only  to  be  successfully  en- 
countered by  superabundant  strength  in  every  portion  of  the 
apparatus.  This  is  well  exemplified  in  the  breakers  turned  out 
from  the  foundries  of  those  manufacturers  who  have  long  made  a 
study  of  this  particular  business,  and  who  have  gradually  added 


an  inch  of  metal  here  and  a  half  inch  there,  as  time  and  trials 
have  developed  the  weak  points  of  the  machine,  until  it  may  ap- 
pear bulky  and  clumsy  beside  the  light  and  elegant  models  of  some 
of  their  later  competitors. 

As  the  ore  usually  passes  directly  from  the  breaker  to  the  rolls 
—better  with  the  interpolation  of  a  short  screen  to  remove  such  as 
is  already  sufficiently  fine;  and,as  in  fine  crushing  the  capacity  of 
the  breaker,  even  when  set  up  to  its  closest  practicable  limits, 
usually  greatly  exceeds  that  of  the  rolls,  a  decided  increase  in  the 
work  performed  can  be  most  economically  and  easily  effected  by 
introducing  a  second  fine  breaker  between  the  coarse  crusher  and 
the  final  pulverizer.  This  machine  may  be  of  quite  light  con- 
struction, should  have  a  very  long,  narrow  jaw  opening — say  2  by 
18  inches — a  slight  "throw,"  and  move  at  a  high  speed. 


The  apparatus  best  suited  for  this  purpose  may  be  brought 
nnder  the  following  heads: 

Stamps.  Multiple-jaw  crushers. 

Ball  pulverizers.  Cornish  rolls. 

Chilian  mills. 

Stamps,  although  universally  known  and  always  reliable,  pro- 
duce far  too  great  a  proportion  of  fine  dust,  besides  being  unnec- 
essarily expensive,  both  as  regards  first  cost  and  subsequent  running. 

The  Ball  pulverizer,  when  properly  constructed,  has  the  merit 
of  compactness,  slight  cost,  ecop.omy  in  running,  and  several  other 
advantages,  but  is  of  insuffi'dent  capacity,  and,  like  stamps, 
is  better  calculated  for  the  production  of  fine  pulp  than  of  the 
material  required  for  calcination. 

Chilian  mills  are  rapid,  economical,  and  effective  pulverizers, 
but  have  usually  had  serious  faults  in  their  construction.  After 
much  experimenting,  Fraser  »&  Chalmers  have  designed  a  mill  of 
this  pattern  that  is  such  a  radical  improvement  on  anything  I 
have  ever  seen  before,  as  to  merit  the  attention  of  every  metallur- 
gist. But  for  calcination,  the  I .'hili  mill  is  not  well  adapted,  as  it 
tends  to  produce  a  fine  powder  rather  than  the  minute  granules 
that  we  prefer  for  calcination. 

Multiple-jaio  Crushers. — The  Blake  multiple-jaw  crusher  em- 
braces series  of  sliding  jaws  actuated  by  pitman  and  toggles,  as  in 
the  ordinary  Blake  crusher.  It  offers  the  advantages  of  the  jaw 
system  of  crushing,  and  can  be  used  in  crushing  material  as  fine  as 

GW  &  Mechanical  Engineer. 



can  ever  be  required  for  ordinary  calcination.  In  crushing  lump 
ore  down  to  a  grain  of  3  mm.  (^  inch),  the  following  set  of  crushers 
might  be  used : 

One  7  by  10  ordinary  Blake  breaker. 

One  3-jawed  2  by  20  multiple-jaw  crusher. 

One  7-jawed  -J  by  24  multiple-jaw  crusher. 

The  portion  that  is  already  crushed  sufficiently  fine  is  removed 
between  the  second  and  third  crushers,  by  means  of  a  revolving 
screen,  the  product  of  the  last  crusher  also  being  elevated  to  the 
same  screen.  The  extraordinary  crushing  surface  obtained  by 
thus  multiplying  the  jaws  is  very  apparent.  Thus,  No.  2  crusher 
has  the  equivalent  of  a  jaw  GO  inches  long,  and  No.  3  of  a  jaw 
168  inches  long. 

A  saving  in  first  cost,  in  power,  and  in  dust  production,  are  some 
of  the  most  important  advantages  claimed  for  this  system  of 

Cornish  Rolls. — Few  machines  can  compare  with  the  Cornish 
roll  for  capacity,  economy,  and  certainty  in  crushing  every  variety 
of  ore  and  matte  for  the  purpose  just  indicated.  But  inasmuch 
as  the  various  patterns  of  this  machine  difi;er  almost  as  much 
among  themselves  in  efficiency  and  capacity  as  they  do  from  the 
other  pulverizers  already  mentioned,  and  as  an  examination  of  a 
large  proportion  of  the  roller  plants  in  actual  use  at  the  present 
time  in  this  country  indicates  a  great  want  of  care  in  both  con- 
struction and  management,  and  a  tendency  to  be  satisfied  with  a 
considerably  lower  standard  of  excellence  than  might  eiisily  be 
attained,  it  seems  desirable  to  draw  attention  to  such  points  as 
seem  to  particularly  demand  supervision  or  reformation. 

Eolls  should  bo  ordered  only  from  the  best  makers,  who  can 
refer  to  numerous  similar  machines  of  their  manufacture  in  long 
and  successful  operation,  nor  should  the  metallurgical  engineer 
forget  that  much  of  the  work  for  which  rolls  are  made,  and  in  the 
performance  of  which  they  give  perfectly  satisfactory  results,  is 
for-  phosphates,  gypsum,  lead  ore,  or  similar  soft  or  brittle  sub- 
stances, whose  crushing  bears  no  relation  to  that  of  the  low-grade 
matte  and  tough  quartzose — or  hard  pyritic— ores  that  are  gener- 
ally the  object  of  calcination.  Certain  low  grades  of  matte,  espe- 
cially when  produced  in  blast-furnaces,  contain  a  large  proportion 
of  various  indefinite  compounds  of  copper,  iron,  and  sulphur  that 
are  almost  malleable,  and  would  inevitably  destroy  any  of  the  ordi- 
nary light-weight,  low-priced  rolls  so  frequently  considered  suffi- 


cient  for  general  purposes,  and  occasionally  placed  in  metallurgical 
establishments  with  mistaken  notions  of  economy. 

A  volume  might  be  written  on  the  subject  of  Cornish  rolls.  I 
must  confine  myself  to  a  glance  at  the  three  most  important  points 
connected  with  their  construction  and  management. 

(a)  Gearing  and  speed. 

(b)  Springs. 

(c)  Shells,  or  tires. 

(a)  Geared  rolls  are  preferred,  by  some  engineers,  for  coarse 
crushing.  For  finer  work,  they  cannot  comiiare  with  rolls  driven 
direct  with  belts.  Geared  rolls  should  have  a  speed  of  about  40 
revolutions  per  minute  A  higher  speed  increases  the  production 
of  dust.  The  fine  rolls  may  be  speeded  100  to  160  revolutions, 
this  speed  being  rendered  practicable  by  direct  belting  and  heavy 
band  wheels  acting  as  fly  wheels.  The  peripheral  velocity  may 
vary  between  GOO  and  1,000  feet  per  minute. 

The  old  system  of  building  rolls  too  weak  for  the  work  they 
have  to  do,  furnishing  insufficient  power  to  drive  them,  and  then 
allowing  them  to  spread  apart  and  shirk  every  hard  lump  that 
happens  to  come  between  them,  need  scarcely  be  considered.* 

(b)  Springs  of  rubber,  or  preferably,  steel,  must  be  used,  or  a 
weak  point  must  be  provided,  that  will  break  when  any  dangerous 
strain  is  brought  upon  the  rolls.  For  coarse  rolls,  I  prefer  extra 
strong  steel  springs,  while  breaking-cups  are  best  adapted  to  fine 
rolls.  Ordinary  car-springs  are  not  stiff  enough  for  the  coarse 
rolls,  when  running  on  hard,  tough  rock,  the  resistance  desired 
under  such  conditions  being  some  50  tons  or  more.  A  good  way 
is  to  group  a  number  of  such  springs  between  two  plates,  and  thus 
form  a  practically  rigid  block  that  bears  against  the  movable  roll 
in  such  a  manner  that  its  elasticity  will  not  come  into  play  until 
the  predestined  compression  limit  is  reached.  By  a  familiar 
arrangement  of  equalizing-levers,  we  distribute  the  strains  uni- 
formly over  both  bearings,  that  the  yielding  roll  may  not  be  forced 
into  an  oblique  position. 

Unless  rolls  are  specially  constructed  for  the  purpose,  nothing 
is  gained  in  setting  them  so  that  their  surfaces  are  in  direct  con- 
tact, even  for  the  finest  crushing,  as  they  will  constantly  choke 
j»nd  give  trouble,  without  yielding  nearly  as  large  an  amount  of 

*  I  regret  to  say  that  tbis  practice  may  be  seen  to-day  in  its  fullest  rievelop- 
ment  in  the  new  concentrator  at  the  Elisabeth  shaft  of  the  Hiuinielfahrt  mine 
0,1  Freiberg. 


product  of  the  desired  fineness  as  when  they  are  set  slightly  apart, 
and  the  product  that  is  not  fine  enough  to  pass  the  screen  is 
returned  to  them. 

(c)  Shells,  or  tires,  may  be  made  either  of  chilled  iron  or  of 
hammered  steel.  The  chilled  tires  are  so  brittle,  and  the  chilled 
surface  frequently  so  unequal  in  quality  and  depth,  that  they  often 
cause  much  annoyance.  It  is  also  a  very  tedious  job  to  dress  them 
into  shape  when  they  require  it. 

Hammered  steel  will  usually  prove  the  more  satisfactory  mate- 
rial, and  can  easily  be  turned  true  when  worn.  The  life  of  the 
shells  will  depend  largely  upon  the  man  in  charge  of  them.  By 
so  distributing  the  stream  of  ore  as  to  throw  the  maximum  of 
work  on  the  least  worn  portions  of  the  tires,  their  axisteuce  can  be 
greatly  prolonged.  I  have  crushed  some  38,000  tons  of  hard  ore 
with  one  set  of  such  tires,  and  am  aware  that  this  duty  has  been 
much  exceeded. 

Finally,  rolls,  like  all  crushing  machinery,  will  work  with  econ- 
omy and  satisfaction  only  when  their  capacity  materially  exceeds 
the  duty  that  is  put  upon  them. 

The  elevator  is  a  necessary  evil,  but  its  delays  and  annoyances 
can  be  greatly  reduced  by  constructing  it  of  a  capacity  far  beyond 
the  requirements  of  the  case.  Chain  elevators  are  not  a  success, 
having  too  many  wearing  parts.  For  heavy  work  I  prefer  to  use 
a  12-inch  six-ply  rubber  belt,  with  heavy,  10-inch  steel  buckets 
riveted  to  the  belt.  A  speed  of  about  250  feet  per  minute  is  satis- 
factory. Wherever  the  inclination  is  not  too  great,  conveyer  belts 
running  over  concave  idler-pulleys,  are  the  most  economical  and 

Perhaps  the  most  useful  and  durable  screens  are  steel  plates 
punched  with  diagonal  holes  and  set  in  a  hexagonal  frame.  The 
plates  can  be  easily  renewed,  and  the  rate  of  screening  can  be 
varied  by  changing  the  level  of  the  lower  bearing.  Screens  are 
seldom  designed  of  sufficient  capacity. 


THE    ROASTING     OF    ORES    IN     LUMP    FORM. 

The  roasting  of  sulphureted  ores  or  copper  in  niounda  or  heaps 
dates  back  beyond  the  age  of  history,  and_,iu  its  most  primitive 
form, is  still  practised  among  barbarous  nations  who  have  evidently 
never  held  communication  with  each  other.  It  is  not  difficult  to 
imagine  its  origin  in  the  midst  of  some  rude  people,  whose  posses- 
sion of  superficial  deposits  of  oxides  and  carbonates  of  copper  had 
taught  them  the  value  of  that  metal  as  obtained  by  a  simple  process 
of  fusion,  while  the  sulphide  ores  that  were  doubtless  encountered 
at  a  slightly  greater  depth  were  thrown  aside  in  heaps  as  worthless 
until  the  spontaneous  combustion  of  some  of  these  wastcpiles, 
brought  about  by  the  decomposition  of  the  sulphides,  and  the 
interesting  discovery  that  ores,  hitherto  considered  valueless, 
would,  after  a  simple  burning,  also  yield  the  coveted  metal,  led 
some  metallurgist  of  that  day  to  the  idea  of  calling  in  the  aid  of 
artificial  combustion  to  hasten  matters.  Nor  has  this  rude  and 
simple  process  undergone  that  general  improvement  that  one  might 
have  expected  when  considering  the  tremendous  advances  made 
in  other  appliances  for  accomplishing  the  same  purposes.  A  some- 
Avhat  careful  inspection  of  nearly  all  the  localities  in  the  United 
States  where  heap-roasting  is  practised  reveals  the  fact  that  the 
results  obtained  are  far  from  satisfactory  in  the  greater  number  of 
instances.  The  amount  of  fuel  employed  and  the  height  and  size 
of  the  heap  are  not  correctly  proportioned  to  the  sulphur  contents 
of  the  particular  ore  under  treatment.  Fragments  of  rock  far 
exceeding  in  size  the  extreme  proper  limit,  as  determined  by  expe- 
rience, are  mixed  with  material  so  fine  as  to  be  fitted  only  for  the 
covering  layer,  and  these  are  dumped  upon  the  ill-arranged  bed  of 
fuel  without  regard  to  the  final  shape  of  the  structure  or  the 
establishment  and  maintenance  of  the  requisite  draught.  Also,  a 
sufficient  quantity  of  proper  material  for  the  all-important  cover- 

THE    ROASTING    OF   ORES    IN    LUMP   FORM.  105 

iug  layer  is  not  applied.  The  result  of  these,  and  some  other, 
deficiencies  is  that  a  small  proportion  only  of  the  ore  is  exi30sed  to 
a  proper  degree  of  heat,  and  the  remainder  of  the  heap  is  pretty 
equally  made  up  of  half-molten  masses  of  clinkers  from  the  inte- 
rior, and  comparatively  raw  and  unburned  material  from  the  outer 
layer.  With  the  exception  of  what  little  sulphur  may  have  been 
driven  off  by  volatilization,  the  ore  after  such  a  calcination  is 
scarcely  better  fitted  for  the  fusion  that  is  to  follow  than  if  it  had 
not  been  roasted.  The  evil  results  of  an  imperfect  preliminary 
calcination  can  only  be  fully  appreciated  after  the  ore  has  passed 
to  the  next  stage  of  treatment;  in  fact,  they  are  so  far-reaching 
that  it  is  impossible  to  express  the  full  measure  of  the  damage  in 
exact  figures.  A  discussion  of  the  effect  of  imperfect  calcination 
and  of  its  remedies  vvill  be  found  under  the  head  of  "Smelting 
Sulphide  Ores  in  Blast  Furnaces."  The  vital  importance  of  the 
process,  and  the  almost  universal  want  of  care  and  supervision  in 
the  carrying  out  of  its  details,  will  justify  this  urgent  remonstrance 
against  its  improper  execution.  Moreover,  the  cost  of  roasting 
properly  is  no  greater  than  that  of  doing  it  imperfectly. 

The  responsibility  of  selecting  heap-roasting  in  contradistinction 
to  the  other  methods  enumerated  for  the  desulphurization  of  an 
ore  must  rest  upon  the  metallurgist  in  charge  of  the  works,  and  is 
a  question  deserving  the  most  careful  consideration;  nor  are  the 
reasons  for  or  against  its  adoption  in  most  cases  so  clear  and  self- 
evident  that  plain  and  unvarying  rules  can  be  laid  down  for  his 
guidance.  In  this,  as  in  many  other  instances,  there  are  usually 
strong  metallurgical,  commercial,  and  sanitary  arguments  that 
should  be  carefully  weighed.  The  contiguity  of  cultivated  land, 
or  even  of  valuable  forests,  would  forbid  the  employment  of  heap- 
roasting  unless  the  arguments  for  its  adoption  were  sufficiently 
powerful  to  outweigh  the  annoyance  of  constant  remonstrances  on 
the  part  of  the  land-owners,  accompanied  by  claims  for  heavy 
damages  from  the  effect  of  the  sulphurous  gases.  For  legal  rea- 
sous,  as  well  as  for  various  other  prudential  and  sanitary  motives, 
it  is  important  to  learn  how  this  damage  is  effected,  and  to  what 
distance  its  ravages  may  extend. 

1.  The  damage  is  caused  solely  by  sulphurous  and  sulphuric 
acids,  neither  arsenical  nor  antimonial  fumes  nor  the  thick  clouds 
of  smoke  evolved  from  bituminous  coal  having  any  appreciable 

2.  The  most  injurious  effects  are  visible  on    young,   growing 


plants;  and  the  more  tender  and  succulent  their  nature,  the  more 
rapid  and  fatal  are  these. 

3.  A  moist  condition  of  the  atmosphere  greatly  heightens  the 
injurious  effects  of  the  gases,  and  as  our  most  frequent  rains  occur 
in  the  spring,  at  the  very  period  during  which  the  crops  and  forests 
are  in  young,  green  leaf,  more  damage  may  be  effected  in  a  few 
days  at  this  season  than  duriug  tlie  entire  remainder  of  the  year. 
The  author  has  seen  a  passing  cloud,  while  floating  over  a  dozen 
active  roast  piles,  absorb  the  sulphurous  smoke  as  rapidly  as  it 
arose,  and,  after  being  wafted  to  a  distance  of  some  eight  miles  by 
a  gentle  breeze,  fall  in  tlie  shape  of  an  acrid  and  blighting  rain 
upon  a  field  of  young  Indian  corn,  withering  and  curling  up  every 
green  leaf  in  the  whole  tract  of  many  acres  in  less  than  an  hour, 

4.  As  might  be  expected,  the  vegetation  nearest  the  spot  where 
the  fumes  are  generated  suffers  the  most,  and  the  direction  of  the 
prevailing  winds,  in  a  fertile  district,  can  be  plainly  determined 
by  the  sterile  appearance  of  the  tract  over  which  they  blow. 

The  most  elaborate  means  for  obviating  this  evil  have  been  tried 
at  the  great  metallurgical  establishments  of  Europe,  and  vast  sums 
have  been  expended  in  this  direction.  The  plans  pursued  in  Eng- 
land tend  more  toward  the  mechanical  deposition  of  the  offending 
substances  in  long  flues  and  passages  (the  first  experimenters  evi- 
dently having  failed  to  realize  that  the  sulphurous  vapors  alone 
caused  the  damage)  while  in  Germany, the  more  scientifically  cor- 
rect method  of  effecting  condensation  and  absorption  of  the  gases 
hy  means  of  various  liquids  and  chemicals  was  pursned.but  with 
scarcely  better  results.  In  the  former  case,  it  was  soon  discovered 
that,  while  the  oxides  of  zinc,  lead,  arsenic,  antimony,  and  various 
other  substances  carried  over  mechanically,  or  as  gases  by  the 
draught,  were  condensed  and  deposited  so  completely  in  the  canals 
that  the  air  issuing  from  the  top  of  the  tall  chimney  was  practically 
free  from  them,  the  percentage  of  sulphurous  and  sulphuric  acids, 
which  alone  are  responsible  for  damage  to  vegetation,  was  not  sen- 
sibly diminished.  Similar  efforts  in  Germany  for  the  absorption 
of  the  sulphur  gases  were  carried  out  with  such  imperfect  and 
ill-adapted  apparatus,  and  on  so  inadequate  a  scale,  that  the  abso- 
lute impossibility  of  a  successful  issue  must  be  apparent  to  any 
one  reading  the  pamphlet  issued  by  the  Freiberg  officials  intrusted 
by  government  with  the  execution  of  the  experiments.  But  how- 
ever insufficient  the  apparatus,  the  results  arrived  at  decisively 
indicated  the  imp  Nsibility  of  disposing  of  the  offending  fumes  by 

THE    ROASTING    OF    ORES    IN    LUMP   FOKM.  107 

any  plan  of  condensation  or  chemical  absorption,  except  on  a  small 
scale  and  with  ususually  dilute  gases. 

The  problem  has  long  been  solved  in  Europe,  in  the  only  rational 
and  economical  manner,  by  utilizing  the  hitherto  destructive 
fumes  for  the  manufacture  of  sulphuric  acid.  This  requires,  of 
course,  the  abolition  of  heap-roasting,  and  the  confinement  of  all 
processes  of  calcination  to  such  closed  kilns  and  furnaces  as  may 
be  placed  in  direct  communication  with  the  leaden  acid  chambers. 
The  very  secondary  position  held  by  agriculture  in  those  sections 
of  our  country  that  furnish  the  material  for  the  principal  smelting 
works  has,  up  to  the  preent  time,  obviated  any  necessity  of  dealing 
with  this  question,  though  some  of  the  largest  copper  smelting 
works  in  the  East  have  already  adopted  the  European  solution  of 
the  problem  as  a  matter  of  profit  rather  than  of  necessity. 

lu  the  case  of  smelting  establishments  of  such  capacity  that  not 
more  than  twenty-five  tons  daily  of  sulphur  are  oxidized  and 
poured  into  the  atmosphere,  it  is  probable  that  all  vegetation  out- 
side of  a  circle  of  four  miles  in  diameter  may,  under  ordinary 
circumstances,  be  considered  safe  from  the  effects  of  the  fumes. 

No  harm  to  man  or  beast  has  ever  been  authentically  reported 
as  resulting  from  the  use  as  food  of  an  article  of  vegetable  origin 
that  has  been  exposed  to  the  corrosive  influence  of  such  gases. 
This  is  a  very  important  point,  and  careful  investigation  and  ex- 
periments have  completely  disproved  the  opposing  arguments  so 
often  made  against  smelting  works  in  Germany  by  certain  stock- 

In  laying  out  the  ground  for  roast-piles,  the  first  point  to  con- 
sider is  the  prevailing  direction  of  the  wind,  great  care  being 
taken  that  the  fumes  shall  neither  be  blown  toward  the  works 
themselves,  nor  toward  the  offices  and  dwelling-houses  in  their 
immediate  neighborhood.  Smelting- works  are  frequently  situated 
in  a  valley,  in  which  the  prevailing  winds  naturally  follow  its 
longitudinal  axis.  In  this  case,  a  tract  of  ground  on  one  side  or 
other  of  the  central  depression,  instead  of  in  its  immediate  course, 
should  be  selected.  By  careful  observation,  and  taking  into  con- 
sideration that  the  prevailing  winds  may  differ  at  different  seasons 
of  the  year,  the  roast  heaps  can  generally  be  so  placed  as  to  give 
no  substantial  ground  for  claims  of  damage  to  agriculture.  Care- 
should  also  be  taken  that  the  selected  tract  is  free  from  any  possi- 
ble chance  of  inundation;  that  it  is  either  perfectly  dry,  or  suscep- 
tible of  thorough   drainage;  that  it  is  not  crossed  by  gullies  oi' 


tleijressions  that  may  serve  as  watercourses  for  the  draiuage  of  the 
surrounding  hills  in  case  of  a  heavy  shower;  that  it  is  protected  as 
far  as  possible  from  violent  winds;  that  snow  does  not  drift  on  it 
badly  in  winter,  and  that  it  is  at  least  as  high  as  the  spot  to  which 
the  ore  is  to  be  transported  for  the  ensuing  operation,  or,  if  this  is 
not  feasible,  at  least  as  high  as  the  elevator  which  is  to  raise  it  to 
the  required  level.  If  possible,  it  should  occupy  an  intermediate 
position,  as  regards  grade,  between  the  shed  in  which  the  ore  is 
prepared  for  roasting  and  the  point  at  which  the  calcined  product 
is  to  be  delivered.  A  fall  of  10  feet  for  the  first  step  and  4-^ 
or  more  for  the  second — total  14^  feet — will  render  possible  the 
establishment  of  a  system  of  handling  and  transportation  that  can 
hardly  be  excelled. 

A  detailed  description  of  such  a  model  plant  will  suffice  as  a 
pattern  that  may  be  varied  to  suit  local  conditions,  always' remem- 
bering that,  under  ordinary  American  circumstances,  the  economy 
of  labor  is  one  of  the  first  conditions  to  be  observed,  and  that  the 
saving  of  25  cents  in  handling  a  ton  of  crude  ore  is  equal  to  a 
dollar  or  more  on  the  ton  of  matte,  and  at  least  two  dollars  when 
estimated  on  the  ton  of  copper. 

Assuming  that  the  metallurgist  is  called  upon  to  prepare  a  yard 
for  heap-roasting  of  ample  size  to  contain  a  sufficient  number  of 
piles  to  furnish  from  80  to  100  tons  daily  of  calcined  material, 
without  encroaching  npon  the  partially  burned  ore,  and  that  the 
contour  of  the  ground  permits  the  requisite  fall  in  each  direction 
— as  already  explained — the  following  plan  may  be  advantageously 
adopted : 

Experience  having  demonstrated  that  an  ordinary  pile  40  feet 
long,  24  feet  wide,  and  fi  feet  high  will  contain  about  240  tons, 
and  burn  for  70  days,  to  which  should  be  added  10  days  for  re- 
moving and  rebuilding,  it  follows  that  each  pile  Avill  supply  ^y 
equal  to  3  tons  of  roasted  ore  daily;  so  that  35  heaps  will  be 
needed  to  furnish  the  full  amount  of  100  tons  daily.  Allowing  30 
feet  for  the  width  of  each  structure,  and  60  feet  for  the  length, 
in  order  to  give  ample  room  for  various  purposes  that  will  be  ex- 
plained hereafter,  an  area  of  75, GOO  square  feet  will  be  required. 

The  frost  being  out  of  the  ground  and  the  surface  dry,  a  rectan- 
gular area  of  the  extent  just  computed  should  be  prepared  by 
means  of  plow  and  scraper,  being  leveled  to  a  perfect  plane,  and 
having  a  slight  slope  toward  one  longitudinal  edge,  or  from  a  cen- 
tral ridse  toward  either  side.     The  black  surface  soil  should   be 

THE    ROASTING    OF   ORES   IN   LUMP    FORM.  109 

removed,  togetner  with  all  sods,  stumps,  and  remains  of  vegeta- 
tion, and  the  space  that  it  occupied  replaced  with  broiien  stones, 
slag,  or  coarse  tailings  from  the  concentrator;  or,  best  and  cheap- 
est of  all,  granulated  slag  from  the  blast-furnace.  This  can  be 
easily  obtained  in  any  desired  amount  by  allowing  the  molten 
scoriae  from  the  slag-spout  to  drop  into  a  wooden  trough,  lined 
with  sheet  iron,  placed  with  a  grade  of  one  inch  to  the  foot,  and 
provided  with  a  stream  of  water  running  through  it,  equal  to  at 
least  sixty  gallons  a  minute.  If  sufficient  fall  is  available,  the 
granulated  slag — graduated  to  any  desired  size  by  the  height 
through  which  it  falls,  velocity  and  amount  of  water,  and  various 
other  trifling  factors  easily  ascertained  by  trial — is  discharged 
directly  from  the  launder  into  dump-carts,  the  water  being  drawn 
off  by  substituting  a  sieve  of  ten  meshes  to  the  linear  inch  for  the 
lower  eighteen  inches  of  the  wooden  trough  bottom.  By  this  sim- 
ple means,  the  best  kind  of  filling  can  be  prepared  and  delivered  at 
the  roastiug-yard  very  cheaply,  the  expense  of  transportation  hardly 
equaling  the  wages  of  the  ordinary  slag-men,  who  may  be  employed 
in  attending  to  the  loading  of  the  carts  and  the  leveling  of  the 
material  when  dumped.  The  entire  area  of  the  rectangle  being 
raised  at  least  two  inches  above  the  surrounding  ground,  a  proper 
surface  is  formed  by  spreading  upon  the  foundation  already  de- 
scribed a  sufficient  quantity  of  clayey  loam.  This  should  be  rolled 
several  times  with  a  heavy  roller  drawn  by  horses,  the  surface 
being  slightly  dampened  from  time  to  time,  until  the  entire  area 
is  as  level  and  nearly  as  hard  as  a  macadamized  road. 

Unless  the  climate  is  an  unusually  dry  one,  and  the  district  free 
from  snow,  it  will  be  better  to  use  gravel  instead"  of  the  loam,  put- 
ting down  a  layer  some  four  inches  thick  over  the  entire  surface  of 
the  roast-yard.  This  will  prevent  mud,  and  the  great  loss  arising 
from  the  treading  of  the  fine  ore  into  the  same. 

If  the  roast-yard  is  to  be  a  permanency,  and  one  is  desirous  of 
obtaining  the  best  results  with  the  least  loss,  a  final  covering  of 
ore-fines  should  bo  added,  the  gravel  being  covered  three  or  four 
inches  deep  with  low-grade  fines.  Nor  should  this  covering  be 
confined  merely  to  the  portion  of  the  ground  that  is  to  be  occupied 
by  the  ore-heaps,  but  should  be  applied  to  the  entire  surface,  in- 
cluding spaces  between  the  heaps,  passageways  at  ends  of  heaps, 
etc.,  etc.  By  so  doing,  there  will  always  be  a  caked  coating  of 
ore-fines  to  shovel  on,  and  the  danger  of  getting  dirt  and  gravel 
mixed  with  the  roasted  ore  will  be  avoided  completely. 


As  the  layer  of  fines  beneath  the  heaps  becomes  gradually  roasted 
through,  it  should  be  removed  with  the  coarse  ore  and  sent  to  the 
furnaces,  its  place  being  supplied  by  fresh  fines  of  the  richest  de- 
scription, for  nowhere  can  fine  ore  be  roasted  so  free  from  any  pos- 
sibility of  loss  as  when  safely  buried  beneath  the  heap. 

Nothing  is  more  important  about  a  roast-yard  than  a  proper 
drainage  system.  If  possible,  the  entire  ground  should  slope 
slightly  toward  the  lateral  lower  track  on  which  the  roasted  ore  is 
removed  to  the  furnaces;  and  where  such  a  gentle  slope  can  be 
obtained,  the  drainage  problem  is  rendered  very  simple  and  per- 
fect; for  a  deep  ditch  run  all  along  the  upper  edge  of  the  mound, 
parallel  with  the  track  just  referred  to,  will  cut  ofE  all  the  surface 
water  from  the  ground  beyond,  and  leave  to  deal  with  only  the 
small  amount  of  water  that  falls  on  the  roast-yard  itself.  This 
water  is  best  removed  by  tile  drains,  laid  underground,  with  fre- 
quent openings  at  suitable  places,  where  there  is  no  danger  of  fine 
ore  being  washed  into  the  drain. 

They  will,  of  course,  have  their  discharge  through  the  bank-wall 
into  the  ditch  that  runs  between  the  lower  track  and  the  bank- 
wall.  Assuming  a  fall  of  some  ten  feet  between  the  spalling-shed 
and  the  ground  under  consideration,  an  elevated  track  is  con- 
structed over  the  central  longitudinal  axis  of  this  rectangle  for  the 
purpose  of  delivering  the  broken  ore  upon  the  heaps.  Where  no 
side-hill  is  available  the  ore  is  carried  up  on  to  the  heaps  in  wheel- 
barrows. The  trestles  to  support  the  track  may  consist  of  sets  or 
bents  of  two  8-inch  by  12-inch  posts  with  8-iuch  by  10-inch  caps 
six  feet  long.  Bents  36  feet  apart  and  properly  braced.  The 
posts  should  be  about  six  feet  apart  at  the  bottom  and  two  or  three 
feet  apart  at  the  top. 

These  bents  support  the  trussed  beams  10  inches  by  12  inches, 
on  edge,  which  carry  the  track  as  shown  in  the  accompanying 
sketch.  (See  Fig.  18).  These  girders  may  be  made  up  of  2-inch 
or  3-inch  planks  spiked  together. 

A  fall  of  an  inch  in  12  feet  will  greatly  facilitate  the  handling 
of  the  loaded  car,  and  offer  little  obstruction  to  the  return  of  the 
empty  one.  The  track  should,  if  possible,  consist  of  T-rails,  12 
pounds  to  the  yard,  firmly  spiked  to  the  longitudinal  stringers,  no 
sleepers  being  necessary;  and  well  connected  with  each  other  by 
fish-plates,  having  two  half-inch  bolts  at  each  end  of  each  rail. 
All  tracks  throughout  the  entire  establishment  should  have  the 
same  gauge;  22  inches  is  a  convenient  standard. 




An  iron-bodied  end-d limping  car,  so  made  as  to  dump  at  right 
angles  to  the  track,  should  be  used.  As  the  heaps  are  some  40 
feet  in  length,  the  area  over  which  the  ore  can  be  distributed  by 
dumping  from  the  car  is  far  too  contracted,  and  the  following 
simple  contrivance  will  be  found  to  save  many  thousand  dollars 
annually  that  would  otherwise  be  expended  in  spreading  the  ore 
bv  hand;  a  plate  of  f-inch  boiler  iron,  30  inches  square,  fitted 
with  a  pair  of  short,  low  rails,  on  three  sides  of  it,  is  so  cut  and 
placed  upon  the  stationary  track  that  the  loaded  car,  striking  first 
the  flattened  extremities  of  one  set  of  the  short  rail  pieces,  while 
the  flanges  of  the  wheels  run  in  corresponding  slits  until  elevated 
upon  the  turntable  by  the  gradually  increasing  height  of  the  short 
rails  referred*  to,  the  heavy  car  may  be  easily  turned  upon  the 
greased  plate  by  a  single  workman,  being  held  and  guided  to  the 
similar  pair  of  short  rails  placed  at  right  angles  to  those  already 
described  by  a  circular  guard  rail,  fastened  at  that  end  of  the  plate 
opposite  to  the  point  of  entrance.  A  temporary  track,  formed  of 
a  pair  of  heavy  rails,  held  firmly  together,  prevented  from  spread- 
ing bv  crossties,  and  supported  by  movable  trestles,  is  laid  at 
right  angles  to  the  main  railroad,  corresponding  exactly  to  a  pair 
of  the  short  side  rails  on  the  turntable  plate.  It  will  be  readily 
seen  that,  by  this  simple  contrivance,  the  extreme  end  of  the  longest 
roast-pile  can  be  reached  with  the  loaded  car,  while  the  turntable 
plate  can  be  shifted  backward  and  forward  until  every  squpre  foot 
of  the  heap  has  received  its  proper  quota  of  ore.  The  accompany- 
ing dimensioned  drawing  illustrates  sufficiently  the  principal 
arrangements  described  in  the  preceding  pages.  If  the  contour  of 
the  surface  permit,  one  longitudinal  side  of  the  prepared  ^avd 
should  be  bounded  by  a  wall  about  four  feet  in  height,  the  top  of 
the  same  being  level  with  the  ground  on  whicii  the  roast-heaps  are 
built,  while  a  railroad  leading  to  the  furnaces  is  constructed  par- 
allel with  it,  in  such  a  manner  that  the  calcined  ore  may  be  wheeled 
on  a  plank  and  dumped  directly  into  cars  without  having  to 
ascend  any  grade,  thus  greatly  lessening  the  expense  of  loading. 
The  labor  and  cost  of  preparing  a  plant,  such  as  has  been  just 
described,  will  be  quickly  repaid  by  the  consequent  avoidance  of 
the  waste  inseparable  from  a  moist  and  muddy  roasting-yard,  and 
especially  from  water  flowing  between  the  heaps.  A  case  came 
under  the  author's  observation,  where  the  want  of  proper  facilities 
for  carrving  off  surface  water  caused  a  loss  estimated  at  ^12,000 
•within  an  hour,  merely  from  the  material  washed  away  by  tin- 

THE    ROASTING    OF    OKES    IX    LUMP    FOKM.  Il3 

back-water  from  a  swollen  ditch,  which  passed  between  the  roast- 
heaps,  but  which,  from  motives  of  economy,  had  been  made  too 
small  to  carry  off  unusual  floods. 

The  height  of  the  pile  must  depend  entirely  upon  the  character 
of  the  ore  aud  the  time  for  calcination  at  the  disposal  of  the  metal- 
lurgist. The  iiigher  the  heap  the  more  fiercely  it  will  heat,  and 
the  longer  it  will  take  to  complete  the  operation.  Consequently, 
where  the  ore  is  rich  in  sulphur,  and  when  time  is  an  object,  as 
where  the  supply  for  the  furnaces  is  small,  heaps  should  be  made 

An  ore  with  15  per  ceut.  sulphur,  which  is,  perhaps,  as  low  as 
can  be  thoroughly  roasted  in  heaps  witliout  the  intermixing  of  a 
considerable  quantity  of  fuel  throughout  with  the  i'cck,  may  be 
piled  up  to  a  height  of  9  feet  advantageously,  while  solid  pyrites 
with  a  sulphur  tenor  of  from  35  to  10  per  cent,  should  never  be 
allowed  to  exceed  G  or  8  feet,  the  measurement  including  only  the 
ore,  and  not  the  layer  of  wood  on  which  it  rests.  The  best  average 
height  for  ordinary  ore  is  7  feet,  under  which  circumstances  it 
will  burn  75  days;  the  time  being  coriespondingly  diminished  or 
increased  by  10  days,  if  C  inches  be  taken  from,  or  added  to,  the 
above  figures.  The  length  of  the  heap  has  little  influence  on  this 
time.  The  following  table  gives  the  result  of  the  roasting  of  large 
quantities  of  various  ores.  In  most  of  these  cases,  frequent  sul- 
phur assays  were  made  of  the  ore  under  treatment;  but  in  a  few 
instances  the  sulphur  was  estimated  from  a  general  knowledge  of 
the  material.  The  heaps  were  thoroughly  covered  and  carefully 
watched,  and  the  combustion  was  kept  at  the  lowest  point  com- 
patible with  safety,  the  sole  object  being  to  obtain  the  most  thor- 
ough possible  roast,  regardless  of  time  or  trouble. 

This  should  be  the  universal  practice;  for  although  the  grade 
of  metal  to  be  produced  in  the  subsequent  fusion  may  not  demand 
such  a  thorough  calcination,  it  is  better  to  roast  a  certain  portion 
of  the  stock  thoroughly,  and  then  reduce,  or  dilute,  the  matte  lo 
the  required  standard  by  the  addition  of  raw  ore.  This  lessens 
expenses  in  various  ways.  It  costs  little  or  no  more  to  roast  an  oie 
thoroughly  than  to  do  so  partially;  and  the  more  completely  the 
sulphur  is  eliminated  from  the  roasted  ore  the  larger  will  be  the 
propor^'ion  of  raw  ore  that  can  be  used  in  the  charge;  and  conse- 
quently the  less  will  be  the  cost  of  calcining  aud  the  losses  from 
fines  of  roasted  ore.  It  is  also  very  easy  to  keep  the  "pitch"  or 
percentage  of  the  matte  produced  at  a  jiroper  point,  when  thor- 


onghly  oxidized  stock  is  always  at  hand.  These  and  various  other 
reasons  that  could  be  mentioned  are  sufficient  to  refute  the  argu- 
ments of  those  who  consider  the  addition  of  raw  ore  peculiarly  in- 
jurious, and  prefer  an  irajierfect  roasting. 


Height  Oiialitv  of  Orp  Percent.      Percent.  Days  No  of 

in  feet.  quality  oi  ure.  Sulphur.        Coi)per.  Burning.  Sample. 

.5...Pyrite 39                W  54  No.     1 

5. .  .  .Clialcopyriie,  with  little  py- 

rite  in  quartz 18  14.3  41  "       2 

5. . .  .Bornite  and  pyrite 31  21.4  53  "       3 

5i...Same  as  No.  1 39                6*  66  "4 

si...         "      No.  2 18  14.3  50  "5 

^...         "      Xo.  3 31  21.4  65  "6 

6....        "      Xo.  1 39               6i  72  "      7 

6....         ••      Xo.  2 18*  14.3  61  "      8 

6.-..         "      Xo.  3 31  21.4  74  "      9 

7 "      Xo,  1,  much  matted  39*              U  94  "     10 

7....         "      Xo.  3 31  21.4  86  "     11 

7i. .   Copper  glance  and  pyrite  in 

quartz 20*  23.4  54  "      12 

The  area  of  the  heap  is  determined  by  the  position  and  size  of 
the  ground  at  disposal,  and  tiie  convenience  of  delivering  the  ore. 
Its  width  is  limited  by  the  distance  to  which  the  covering  material 
can  be  conveniently  thrown  with  a  shovel,  and  by  the  room  be- 
tween the  bents  that  support  the  track;  24  feet  in  width  by  40  in 
length  is  a  very  convenient  size,  smaller  heaps  demanding  consid- 
erably more  labor  and  fuel  to  the  ton  of  ore.  With  36  feet  between 
the  bents,  an  ample  border  of  6  feet  will  be  left  on  each  side  of 
the  pile  for  collecting  the  tines,  wheeling  the  same  wherever  re- 
quired, and  fully  securing  the  wood-work  against  all  danger  of  fire. 
Risk  from  fire  is  further  obviated  by  elevating  the  foundation  sill 
from  which  the  uprights  arisp,  upon  a  wall  of  slag-brick,  3  feet  or 
more  in  height.  A  pile  of  the  dimensions  referred  to,  24  feet  by 
40  feet  square,  and  6  feet  high,  will  contain  ahout  240  tons  of  ordi- 
nary ore,  and  should  be  built  in  the  following  manner :f 

The  corners  of  the  rectangular  space  on  which  it  is  to  be  erected 
should  be  indicated  by  stakes,  or,  if  the  same  size  is  to  be  perma- 


+  If  the  furnaces  are  not  too  much  pressed  for  ore  it  is  more  economical  to 
still  further  increase  the  size  of  tlie  heaps;  40  by  80  feet,  and  7  feet  high  is 
none  too  large. 

THE    ROASTING    OF    ORES    IN    LUMP   FORM.  115 

mently  retaineci,  by  large  stones,  or  better,  blocks  of  slag,  imbedded 
in  the  ground.  The  sides  of  the  area  being  indicated  by  lines 
drawn  on  the  ground  to  guide  the  workman,  the  entire  space 
should  be  covered  evenly  to  the  depth  of  four  or  six  inches  with 
fine  ore  from  the  spalling-shed.  This  layer  of  sulphides  answers 
several  purposes;  in  the  first  place,  it  prevents  the  baking  and 
adhering  to  the  ground  of  the  coarser  ore,  which,  especially  when 
much  matte  is  formed,  sticks  to  the  clayey  soil  to  such  an  extent 
as  to  tear  up  and  injure  the  foundation,  besides  mixing  worthless 
dirt  with  the  ore,  and  causing  a  loss  of  the  latter  when  attempts 
at  separation  are  made.  It  also  forms  a  distinct  boundary  line 
between  the  worthless  and  valuable  materials,  and,  when  left  un- 
disturbed during  two  or  three  operations,  becomes  itself  so  thor- 
oughly desulphurized  that  the  upper  half  or  more  may  be  scraped 
up  with  shovels  and  added  to  the  roasted  ore,  its  place  being  filled 
by  a  fresh  supply  of  fines.  This  operation  completed,  the  fuel  is 
next  arranged  by  an  experienced  workman  in  a  regular  and  sys- 
tematic manner.  The  quality  and  size  of  the  wood  is  a  matter  of 
«ome  moment,  and  must  be  determined  for  each  individual  case,  it 
being  evident  that  that  variety  of  fuel  that  yields  the  greatest 
amount  of  heat  for  the  longest  time  possesses  the  highest  money 
value,  provided  the  ore  is  of  such  a  nature  as  to  bear  the  tempera- 
ture produced  without  fusing.  As  most  sulphide  ores  will  not 
stand  the  heat  generated  by  a  thick  bed  of  sound,  dry,  hard  wood, 
it  frequently  happens  that  a  cheaper  variety  answers  the  purpose 
better.  The  outside  border  of  wood  that  corresponds  to  the  edges 
of  the  heap  should  be  of  better  quality,  as  no  such  degree  of  heat 
is  attainable  there  as  in  the  interior  of  the  pile.  Therefore  a  large 
proportion  f^t  the  bed  may  be  made  up  of  old  rails,  logs,  gnarled 
and  knotted  trunks  that  have  defied  wedge  and  beetle,  and  such 
sticks  of  cordwood  as  are  daily  thrown  out  from  wood-burning 
boilers  and  calcining-furnaces  as  too  crooked  and  misshapen  to 
enter  a  contracted  fireplace.  Such  miscellaneous  fuel  causes  some- 
what greater  labor  in  arrangement;  but  whatever  the  material,  it 
must  be  placed  with  such  care  and  skill  as  to  form  a  solid  and 
sufficient  bed,  varying  in  depth  from  4  to  10  inches  according  to 
the  behavior  of  the  ore.  However  rough  and  irregular  the  greater 
portion  of  the  fuel  at  our  disposal  may  be,  enough  cordwood  of 
■even  length  and  diameter  should  be  selected  to  form  a  four-foot 
border  around  the  entire  heap  and  just  within  the  side-lines  of  the 
area;  for  the  even  and  regular  kindling  of  the  heap  depends  con- 


siderably  npon  the  proper  arrangement  of  this  border.  Sticks  of 
cordwood  not  larger  than  5  inches  iu  diameter  should  be  laid  side 
by  side  across  both  ends  and  sides  of  the  area.  Across  this  layer, 
small  wood  is  again  piled  nntil  this  four-foot  border  has  been  built 
np  to  the  height  of  some  10  inches,  brushwood  and  chips  being 
scattered  over  the  surface  to  fill  np  all  interstices,  while  canals  6 
inches  wide,  filled  with  kindlings,  are  formed  at  intervals  of  8  or 
10  feet,  leading  from  the  outer  air  and  communicating  with  the 
chimneys  in  the  center  line  of  the  heap.  The  empty  area  within 
this  encircling  border  is  now  filled  with  the  poorer  quality  of  fuel, 
all  sticks  laid  parallel  and  with  as  much  regularity  as  possible,  to 
cover  all  cracks  and  interstices,  that  no  ore  may  fall  through  the 
wood,  and  to  cover  over  the  draught-canals  in  such  a  manner  that 
thev  shall  be  neither  choked  nor  destroyed  by  the  superincumbent 

The  chimneys,  which  assist  materially  in  rapidly  and  certainly 
kindling  the  entire  heap,  are  formed  of  four  worthless  boards  nailed 
lightly  together  in  such  a  manner  that  two  of  the  opposite  sides 
stand  some  eight  inches  from  the  ground,  thus  leaving  spaces  that 
communicate  with  the  draught-canals  referred  to,  and  toward 
which  several  of  the  latter  converse.  For  a  heap  40  feet  in  length, 
three  such  chimneys,  eight  inches  square,  will  suffice.  'J'hey 
should  project  at  least  two  feet  above  the  proposed  upper  surface 
of  the  structure,  that  no  fragments  of  ore  may  accidentally  enter 
the  flue  opening  and  destroy  its  draught.  In  certain  localities, 
where  even  old  boards  are  too  valuable  to  be  needlessly  sacrificed, 
two  er  three  medium-sized  sticks  of  cordwood  may  be  wired  to- 
gether to  form  the  chimney;  or  old  pieces  of  sheet-iron,  such  as 
condemned  jig-screens,  worn-out  corrugated  roofing-iron,  etc.,  mav 
be  so  bent  and  wired  as  to  form  a  permanent  and  sufficient  passage, 
while  this  material  will  answer  for  several  operations.  The  chim- 
neys being  placed  in  position,  equidistant,  and  on  the  longitudinal 
center  line  of  the  bed  of  fuel,  and  held  upright  bv  temporarv 
wooden  supports,  the  heap  is  ready  to  receive  the  ore.  This  is 
brought  in  carloads  of  1,500  or  2,000  pounds  from  the  spalling- 
shed,  and  weighed  en  route  on  track-scales.  It  is  dumped  on  a 
portable  wooden  platform  about  eight  feet  square,  to  prevent  the 
deranging  of  the  wood  from   the  fall  of  so  heavy  a  mass  of  rock 

*  An  excellent  paper  on  heap  roastiuir  in  Vermont,  by  Mr.  William  Glenn, 
mar  be  found  in  the  Engineering  and  Mining  Journal  for  December  8, 1883. 


from  a  lieight  of  ten  feet  or  thereabout.  The  first  few  carlcads 
are  heaped  about  the  chimneys,  and  the  platform  is  changed  from 
place  to  place  as  convenience  demands,  until  the  bed  of  wood  is 
thoroughly  protected  by  a  thick  layer  of  ore.  The  remainder  of 
the  process  is  a  very  simple  operation.  The  cars  of  ore  are  dumped 
in  turn  over  the  entire  area  by  a  systematic  shifting  of  the  tempo- 
rary pair  of  rails  already  described,  and  the  heap  formed  into  a 
shapely  pyramid,  with  sharp  corners  and  an  angle  of  inclination  of 
some  42  degrees,  or  as  steep  as  the  ore  will  naturally  lie  without 
rolling.  The  main  body  of  the  structure  is  formed  of  the  coarsest 
class  of  ore;  the  ragging  is  next  placed  upon  the  pile,  forming  a 
comparatively  thick  covering  at  the  part  uearest  the  ground,  and 
gradually  thinning  out  toward  the  top  and  on  the  upper  surface. 
Its  thickness  depends  on  the  amount  available,  and  no  fears  need 
be  entertained  of  its  having  an  unfavorable  influence  on  the  calci- 
nation; for,when  carefully  separated  from  the  finest  class,  a  heap 
formed  entirely  of  ragging  will  give  reasonably  good  results.  The 
extreme  outside  edge  of  the  ore,  when  all  is  in  place,  should  not 
entirely  cover  the  external  border  of  wood.  At  least  a  foot  of 
uncovered  fuel  should  project  beyond  the  layer  of  ragging,  both  to 
prevent  the  ore  from  sliding  off  its  bed  as  well  as  to  insure  a  thor- 
ough kindling  of  the  outer  covering  of  mineral.  The  amount  of 
wood  required  properly  to  burn  a  heap  of  240  tons  of  ore  will  vary 
greatly  with  the  composition  of  the  latter,  standing  in  direct  pro- 
portion to  its  sulphur  contents,  and  es^jecially  to  the  amount  of 
bisulphides  present,  but  may,  on  the  average,  be  estimated  at  6 
cords,  or  one  cord  of  wood  to  40  tons  of  ore.  In  smaller  heaps,  this 
proportion  must  be  considerably  increased. 

It  is  the  common  practice  to  use  far  too  much  wood  in  heap- 
roasting.  This  causes  too  great  a  heat  at  the  commencement  of 
the  operation,  and  brings  about  various  irregularities,  such  as  local 
sintering  and  matting  of  the  ore,  with  stoppage  of  air-circulation 
at  these  points,  so  that  when  the  heap  is  finished,  we  find  at  vari- 
ous points  several  square  feet  of  fused  sulphides  on  the  bottom. 
Above  this  comes  a  siliceous  skeleton  of  an  extent  corresponding  to 
the  amount  of  matte  which  has  been  liquated  out  of  it.  And 
above  this  still,  a  large  body  of  unroasted  ore,  which  has  entirely 
escaped  the  firo. 

I  have. never  seen  heap-roasting  more  perfectly  executed  than  at 
the  Spanish  mines  of  TJio  Tinto.  The  material  treated  is  the  rich 
ore  that  is  culled  from  the  ordinary  low-grade  pyrites,  and   is  re 


served  for  smeltins;  iu  blast-furnaces.  It  carries  some  8  or  9  per 
cent,  of  copper,  aiul  is  a  solid  mass  of  irou  aud  copper  pvrites. 

The  most  interesting  features  of  this  hecp-roasting  are  the  very 
small  proportion  of  wood  used,  and  the  unusual  height  of  the  pile. 
It  is  built  very  much  in  the  shape  of  a  circular  haystack,  being 
some  fourteen  feet  high  in  the  center  of  the  cone,  and  about  thirty- 
three  feet  in  diameter.  It  usually  contains  something  over  400 
tons  of  ore,  much  of  it,  near  the  center,  being  in  pieces  the  size  of 
a  child's  head.  Only  two-thirds  of  a  cord  of  wood  is  employed, 
aud  this  is  distributed  among  twelve  fireplaces,  constructed  roughly 
of  rock,  aud  spaced  equidistant  about  the  circumference  of  the 
pile.  They  penetrate  to  a  depth  of  only  four  feet,  so  that  the 
major  portion  of  the  pile  has  no  wood  under  it.  When  the  heap 
is  lighted,  only  the  small  fraction  of  ore  close  to  the  fireplaces  is 
kindled;  and  even  here  the  amount  of  wood  is  so  small  that  the 
heat  is  very  slight  and  evanescent.  From  these  twelve  points  at 
the  circumference,  the  fire  gradually  creeps  toward  the  center, 
while  the  heap  is  thoroughly  covered  with  fines,  and  the  tempera- 
ture kept  far  lower  than  in  ordinary  heap-roasting.  It  consequently 
takes  six  to  nine  months  to  burn  one  of  these  400-tou  piles. 
Bat  the  result  is  a  triumph  of  skill,  scarcely  ever  a  pound  of  matte 
being  formed,  while  in  the  various  heaps  which  were  completed 
and  partially  demolished  for  the  cupolas,  I  was  unable  to  find  even 
a  fragmeut  of  unoxidized,  or  badly  roasted  ore.  With  the  excep- 
tion of  a  few  kernels,  the  lumps  were  oxidized  to  their  very  center. 
I  was  informed  that  any  increase  in  the  amount  of  wood  used  ta 
kindle  the  pile  was  a  drawback  rather  than  an  advantage. 

I  have  no  doubt  that  we  use  far  more  wood  than  is  conducive  to 
wood  roasting,  while  to  attempt  to  hurry  a  process,  at  the  expense 
of  doing  it  properly,  is  certainly  not  profitable.  Conditions  differ 
too  orreatly  to  admit  of  any  hard  and  fast  rules  in  such  matters, 
and  every  metallurgist  must  determine  for  himself  just  how  per- 
fectly it  will  pay  him  to  carry  out  this  process,  aud  how  long  a 
time  he  can  afford  to  spend  in  doing  it.  He  will  probably  arrive 
at  the  conclusion,  if  he  gives  the  subject  proper  attention,  that  it 
will  remunCi'ate  him  handsomely  to  roast  far  more  slowly  and  more 
perfectly  and  with  less  wood  than  he  has  ever  attempted  to  do 
before.  For  the  ore  that  is  tied  up  in  roast-heaps  can  only  be 
charged  with  what  it  has  cost— that  is,  the  expen«P3  of  mining, 
crushing,  and  putting  into  heaps — and  the  interest,.  i»t  10  p^r  cent, 
on  the  cost  of  even  50,000  tons  of  ore  at  x3  per  toii^  is  only  #41  per 

THE    ROASTING    OF    ORES   IN    LUMP    FORM.  119 

day;  a  sum  completely  insignificant  compared  with  the  gain  aris- 
ing from  the  better  grade  of  matte  and  the  lessened  troubles  in 
furnace  management  that  will  result  from  even  a  very  slight 
average  improvement  in  the  roasting  process. 

The  fine  ore  that  is  to  form  the  external  layer,  and  on  which 
depends  largely  the  success  of  the  process,  is  seldom  placed  upon  the 
top  of  the  heap  until  after  it  is  fired.  Perhaps  the  most  judicious 
practice  is  to  cover  the  sides  of  the  pile  with  a  very  thin  layer, 
scattering  it  evenly  with  a  shovel,  and  leaving  the  upper  surface, 
as  well  as  a  space  eighteen  inches  broad  at  the  bottom  uncovered; 
for  if  the  fine  ore  is  thrown  carelessly  upon  the  lower  circumfer- 
ence of  the  pile,  the  draught  is  decidedly  hampered  and  the  fire 
stifled  before  getting  fairly  under  way.  For  an  average  ore,  an 
amount  of  fines  equal  to  10  per  cent,  of  its  total  weight  is  ample; 
of  this,  eight  tons  may  be  strewn  lightly  upon  the  sides  of  the 
heap  as  just  described,  the  remaining  16  tons — assuming  the  en- 
tire contents  to  be  240  tons — being  arranged  in  small  piles  upon 
the  empty  space  between  the  roast-heaps,  where  it  is  easily  acces- 
sible to  the  shovel.  The  lighting  should  be  done  just  as  the  day 
shift  is  quitting  work,  as  the  dense  fumes  of  wood  smolce,  strongly 
saturated  with  pyroligneous  acid  and  the  various  gaseous  com- 
pounds of  sulphur  and  arsenic,  among  which  sulphureted  hydrogen 
is  always  plainly  distinguishable,  are  almost  unbearable. 

If  possible,  fine  weather  should  be  selected  for  this  purpose;  for 
although  no  ordinary  rain  is  capable  of  extinguishing  a  well-lighted 
roast-heap,  it  may  still  interfere  greatly  with  kindling  a  new  one, 
and  is  quite  likely  to  cause  subsequent  irregularities  in  the  course 
of  the  process.  There  are  several  different  methods  of  firing  a  roast- 
heap— such  as  lighting  it  only  on  the  leevvard  side,  and  letting  the 
fire  creep  back  against  the  wind,  kindling  it  through  the  draught- 
chimneys,  etc.,  eacli  of  >which  has  its  advocates  among  roasting 
foremen;  but  long-continued  observation  has  shown  tliat  no  ad- 
vantage is  gained  by  any  of  these  irregular  methods,  and  the  most 
sensible  and  successful  practice  is  to  light  it  as  quickly  and  thor- 
oughly as  possible  by  applying  a  handful  of  cotton  waste,  saturated 
Avith  coal  oil,  or  a  ladle  of  molten  slag,  to  the  kindling-wood  at 
the  mouth  of  each  of  the  draught-canals,  these  being  some  ten  or 
more  in  number,  as  already  described.  As  the  success  of  the 
entire  operation  depends  principally  on  the  management  of  the 
heap  for  the  first  few  days  after  kindling,  it  will  be  necessary  to 
study  somewhat  in  detail  the  phenomena  that  it  should  normally 


exhibit  during  this  critical  period,  always  bearing  in  mind  the 
impossibility  of  laying  down  any  fixed  rules  that  shall  apply  to  all 
circumstances  and  to  every  variety  of  material. 

Under  ordinary  circumstances,  the  heap  may  best  be  left  en- 
tirely to  itself  for  from  four  to  six  hours  after  lighting,  oare  merely 
being  taken  that  the  kindling  burns  freely,  and  that  the  draught- 
holes  conmuiuicatc  with  their  respective  chimneys.  At  the  expi- 
ration of  this  time,  if  the  tire  has  spread  well  over  the  entire  area, 
about  one-half  of  the  remaining  fines  tliat  nave  been  provided  for 
covering  should  be  scattered  lightly  upon  the  heap;  the  lower 
border  and  upper  surface,  which  have  hitherto  been  left  unpro- 
tected, now  receive  a  thin  application,  wnile  the  lateral  coating 
is  rendered  somewhat  thicker  and  more  impervious.  If  matters 
pursue  a  normal  course,  the  6arly  morning — twelve  hours  after 
firing — should  see  the  heap  smoking  strongly  and  equally  from 
innumerable  interstices  produced  by  the  settling  of  the  whole 
mass,  due  to  the  disappearance  of  the  thick  foundation  of  fuel. 
Dense  pillars  of  opaque,  yellow  smoke,  smelling  strongly  of  sul- 
phurous acid,  arise  from  the  site  of  each  chimney;  although  if 
these  were  constructed  of  wood,  no  sign  of  them  will  remain  ex- 
cept a  few  charred  fragments,  resting  in  a  slight  depression,  which 
marks  their  sites.  The  entire  surface  will  be  found  damp  and 
stickv,  and  the  covering  material  will  have  already  formed  quite  a 
perceptible  crust,  from  the  adhesion  of  its  paii^ticles.  This 
"sweating,"  as  it  is  termed,  arises  from  the  distillation  products 
of  the  fuel — owing  to  its  very  miperfect  combustion — and  from 
the  moisture  contained  in  the  ore.  A  yellowish  crust  surrounding 
the  vents  from  which  the  strongest  currents  of  gas  are  seen  to 
issue  indicates  the  presence  of  metallic  sulphur,  the  volatilization 
of  the  first  loosely  bound  atom  of  which  begins  soon  after  the 
wood  is  fairly  lighted.  Its  quantity  depepds  on  the  proportion  of 
bisulphides  in  the  roast,  as  well  as  on  the  freedom  with  which  air 
is  admitted ;  the  scarcity  of  oxygen  and  a  high  temperature  favoring 
its  direct  volatilization,  while  an  abundance  of  air  and  a  moderate 
heat  influence  the  plentiful  generation  of  sulphurous  acid. 

During  this  first  day,  the  newly  kindled  heap  will  require  close 
and  constant  attention  to  prevent  any  undue  local  heating;  nor  is 
it  at  all  uncommon  to  find  that  some  neglected  fissure  has  increased 
the  draught  to  such  an  extent  as  to  cause  the  sintering  or  partial 
fusion  of  several  tons  of  ore  at  that  point.  The  principal  signs 
by  which  the  experienced  eye  judges  of  the  condition  of  atfairs  are 

THE    ROASTING    OF   ORES   IX    LUMP   FORM.  121 

the  color  of  the  gas  and  tlie  rapidity  with  which  it  ascends;  the 
anioinit  of  settling  and  consequent  fissuring  of  the  covering  layer; 
and,  above  all,  the  degree  of  heat  at  diSerent  parts  of  the'surface. 
A  light,  bluish  gas,  nearly  trans^iarent,  and  ascending  in  a  rapid 
current,  is  a  sign  that  the  heat  is  too  great  at  that  point,  and  the 
admission  of  air  too  free.  The  fissuring  of  the  crusted  covering 
material,  after  the  general  and  extensive  sinking  caused  by  the 
consumption  of  the  fuel,  indicates  a  rapid  settling  that  can  only 
arise  from  the  melting  together,  and  consequent  contraction,  of 
the  lumps  of  ore.  All  these  conditions  are  met  by  a  single  remedy ; 
that  is,  covering  the  surface  at  that  point  more  thoroughly  with 
lines,  by  which  means  the  air  is  excluded,  the  rapidity  of  the  oxi- 
dation process  diminished,  and  the  temperature  lowered.  It 
should  not  be  supposed  that,  because  the  interstices  that  exist  in 
the  upper  part  of  the  heap  alone  show  evidences  of  heat  and  gas, 
those  cracks  and  openings  that  have  been  left  nearer  the  ground 
are  of  no  importance;  these  are  the  draught-holes,  while  the 
former  constitute  the  chimneys,  and  it  is  to  the  condition  of  the 
lower  border  of  the  pile  that  our  attention  should  be  most  fre- 
quently directed  in  regulating  the  proper  admission  of  air.  A  few 
shovelfuls  of  Hue  ore  judiciously  applied  at  the  base  of  the  heap 
will  often  have  more  effect  than  a  carload,  scattered  aimlessly 
over  the  surface. 

Only  an  experienced  laborer  can  manage  a  roast-heap  to  the  best 
advantage,  nor  is  it  possible  to  establish  fixed  rules  for  the  guidance 
of  this  process,  varying  conditions  demanding  totally  different 
treatment.  In  a  general  way,  it  may  be  said  that,  after  somewhat 
subduing  the  intense  heat  caused  by  the  sudden  combustion  of  so 
large  an  amount  of  wood,  the  attendant  should  confine  himself  to 
scattering  material  in  a  thin  layer  over  the  sides  and  top  of  the 
structure,  and  effectually  stopping  up  such  holes  and  crevices  as 
seem  to  be  the  vents  for  some  unusually  heated  spot  below. 

Hy  the  third  day  large  quantities  of  sublimated  sulphur  will  be 
found  upon  the  surface,  in  many  places  melting  and  burning  with 
a  blue  flame.  It  is  now  necessary  for  the  attendant  to  ascend  to 
the  top  of  the  heap,  to  properly  examine  the  upper  surface,  and 
place  additional  covering  material  on  such  portions  as  still  seem 
too  hot.  In  doing  this,  a  disagreeable  obstacle  is  encountered  in 
the  clouds  of  sulphurous  gas,  which,  to  one  unaccustomed  to  the 
task,  seem  absolutely  stifling.  By  taking  advantage  of  their  mo- 
mentary dispersion    by  currents  of  air,  and   retreating  when   they 


become  too  thick,  no  difficulty  need  be  experienced  in  covering 
the  npper  surface  of  tlie  heap  as  thoroughly  and  carefully  as  any 
other  part  of  it. 

If  the  process  of  combustion  seems  to  l.uive  spread  equally  to  all 
parts  of  the  pile,  nothing  need  now  be  done  except  daily  to  scatter 
a  few  shovelfuls  of  fines  over  such  heated  spots  as  seem  to  require 
it;  but  if  any  isolated  corner  of  the  heap  has  failed  to  kindle,  or, 
having  once  caught  fire,  has  now  become  cold  and  ceased  to  smoke, 
it  is  necessary  to  draw  the  fire  in  that  direction.  This  can  be 
accomplished  with  ease  and  certainty  by  any  one  accustomed  to 
the  work;  for  there  is  no  danger  of  a  roast-heap  becoming  extin- 
guished when  once  fairly  kindled.  Certain  isolated  spots — espe- 
cially corners  and  angles — may  fail  to  become  properly  ignited, 
but  by  opening  a  few  draught-holes  in  the  neighborhood  the  fire 
will  surely  spread  wherever  unburned.  sulphides  still  exist.  Be- 
ginning at  the  end  of  the  first  week,  and  continuing  for  a  month 
or  more,  a  certain  amount  of  sulphur  may  be  obtained  by  forming^ 
18  or  20  circular,  ladle-shaped  holes  about  14  inclies  in  diameter 
and  7  inches  deep  in  the  upper  surface  of  the  heap,  and  lining 
them  carefully  with  partially  roasted  fine  ore,  so  that  they  may 
retain  the  molten  metalloid.  The  impure  sulphur  may  be  ladled 
out  twice  a  day  into  wooden  molds;  but  the  impurity  of  the  prod- 
uct, caused  by  the  great  quantity  of  ore-dust  and  cinders  constantly 
falling  into  the  melted  material,  and  the  extremely  scant  produc- 
tion of  a  substance  that  is  hardly  worth  saving,  discourages  the 
general  adoption  of  the  practice,  although  at  some  of  the  older 
German  works  it  is  still  kept  up.  Experiments  made  with  the 
greatest  possible  care  saved  only  one-tentli  of  one  per  cent,  of  the 
total  weight  of  the  ore  from  a  ;}(>  per  cent,  bisulphide  ore. 

With  certain  varieties  of  ore,  the  sulphur,  instead  of  collecting 
in  a  concentrated  form  at  the  principal  issuing  vents  of  the  strong- 
est currents  of  gases,  condenses  over  the  entire  surface  in  a  thin 
layer,  and  upon  raelting,cements  and  agglutinates  the  fine  particles 
of  the  covering  layer  in  such  a  manner  as  to  form  an  almost  im- 
permeable envelope  In  such  cases  this  crust  must  be  destroyed, 
from  time  to  time,  with  an  iron  garden-rake,  or  the  process  of 
calcination  may  be  delayed  for  weeks  beyonti  its  customary  limit 
from  the  lack  of  sufficient  oxygen  to  maintain  the  proper  rate  of 
combustion.  If  arsenic  is  present,  even  in  the  smallest  quantities, 
it  will  soon  make  itself  visible  as  beautiful  orange-colored  realgar, 
AsS,  and  minute  clusters  of  white,  glistening  crystals  of  arsenious 


oxide,  wliich  usually  form  at  the  upper  orifices  of  tlie  accideutal 
draught-canals  that  communicate  with  the  interior  of  the  heap. 

A  strong  and  jDersisteut  wind  from  any  one  direction  has  an  un- 
favorable effect  on  the  process  of  heap-roasting,  driving  the  fire 
toward  the  leeward  side,  and  cooling  those  portions  that  feel  the 
direct  influence  of  the  air-current  to  such  an  extent  that  one-fourth 
or  more  of  the  heap  may  remain  in  a  raw  condition.  It  is  a  some- 
what remarkable  fact  that,  while  it  is  almost  impossible  to  quench 
a  roast-heap  with  water,  unless  completely  flooded  for  a  considera- 
ble length  of  time,  a  simple  excess  of  the  very  element  most  favor- 
able to  its  perfect  combustion  should  have  the  power  to  extinguish 
it.  If  this  annoying  circumstance  repeats  itself  with  any  fre- 
quency, it  will  be  necessary  to  erect  a  high  board  fence  on  that  side  of 
the  yard  whence  the  most  persistent  winds  prevail.  Moderate  rain 
and  snow  have  little  influence  on  the  course  of  the  process,  except 
in  so  far  as  they  may  cause  serious  chemical  and  mechanical  losses. 
It  is  only  after  a  heavy  shower  or  sudden  thaw  that  the  great  ad- 
vantage of  numerous  and  well-preserved  ditches  surrounding  the 
entire  area,  and  even  leading  between  the  heaps  theirselves,  is 
fully  realized  and  appreciated.  When  wet  weather  supervenes,, 
after  a  long  period  of  drought,  the  amount  of  copper  dissolved 
from  the  soluble  sulphate  salts  formed  during  the  extended  term 
of  dryness  may  be  so  large  as  to  repay  some  efforts  to  recover  it. 
By  simply  leading  the  drainage  from  the  roast-yard  into  two  old 
brewer's  vats  partially  filled  with  scrap-iron,  during  one  summer,. 
3o,546  pounds  of  40  per  cent,  precipitate  were  collected. 

We  have  already  pointed  out  the  necessity  of  guarding  against 
this  loss  by  every  possible  means  at  our  disposal;  but  even  with 
every  care  a  considerable  loss  from  this  source  cannot  be  avoided  in 
any  ordinary  climate. 

Mr.  Wendt*  gives  some  important  figures  bearing  on  this  point, 
relating  to  heap- roasting  as  formerly  practised  at  Ducktown, 
Tenn.,  where,  however,  the  rainfall  is  exceptionally  great.  We 
quote  also  his  estimates  of  cost,  which,  taking  into  account  the- 
low  cost  of  fuel  and  labor,  correspond  closely  with  our  own. 

"Ore-roasting,  as  thus  carried  out  (in  heaps),  was  a  very  eco- 
nomical process  in  point  of  labor  and  fuel.  On  an  average,  one 
cord  of  wood  was  consumed  for  40  net  tons  of  ore  for  each  fire. 
The  cost  of  labor  in  the  first  fire  was  5  cents  per  1,000  pounds  for 

*  See  The  Pyrites  J)epo.iits  of  the  AUegliaiiies,  by  A.  F.  Wendt,  New  York, 
1866,  page  19. 



both  Mary  and  East  Tennessee  ores;  for  the  second  fire,  7  cents  and 
6  cents  respectively  were  paid ;  and  for  fine  ores,  the  pay  was  12  cents 
per  M. 

"The  exact  cost  per  net  ton  of  ore  was  as  follows: 

A  cord  of  wood  at 
Labor,  1st  fire. . . . 
Labor,  2d  fire  ... . 


Total,  per  ton 10.42 

"The  losses  of  copper  in  the  above-described  roasting  have  been 
very  generallv  ignored  in  judging  of  its  expense.  At  least,  proper 
emphasis  has  never  been  laid  on  them. 

"  Owing  to  an  unexplained  difference  of  several  hundred  thou- 
sand pounds  between  the  fine  copper  produced  at  the  Ducktown 
smelter  during  a  period  extending  over  several  years,  and  the 
monthly  fine  copper  statements  arrived  at  by  deducting  one  and 
one-quarter  unit  from  the  assay  value  of  the  ores  produced,  the 
writer's  attention  was  forcibly  called  to  this  subject.  A  careful 
series  of  experiments  was  instituted;  the  results  v.ere  rather  star- 
tling. Repeated  analysis  of  ore  weighed  into  a  roast-pile,  and 
analysis  and  weighing  of  this  same  ore  when  sent  to  the  matte 
furnaces,  proved  an  almost  incredible  loss. 

"From  the  large  number  of  experiments  and  analyses,  I  quote 
the  following  striking  examples: 

Pile  Xo.  349. — Mary  Ore. 

Gross  Weight  of  Ore. 

Per  Cent.  Water. 

Per  Cent.  Copper. 

Fine  Copper,  Pounds. 









741,667  pounds  raw  ore  contained  36,410  pounds  copper. 

"The  pile  after  roasting  weighed  741,716  pounds— assayed  3.31 
per  cent,  copper — equivalent  to  24,985  pounds  fine  copper;  11,125 
pounds  copper,  or  31.4  per  cent,  of  the  contents  of  the  pile,  had 
been  lost  while  roasting;  170  days  were  consumed  in  roasting  the 
ore  and  09  days  in  removing  it  to  the  smelting-furnaces.  Hence, 
the  ore  lay  exposed  to  the  weather  for  239  days,  that  is,  eight 



Pile  Xo.  447. — Mary  Oke. 

Gross  Weight  of  Ore. 

Per  Cent.  Water. 

Per  Cent.  Copper. 

Fine  Copper,  Pounds. 

























464,505  gross  pounds  ore  contained  21,482  pounds  copper. 

''Weight  of  the  roasted  ore  was  495,566  pouuds,  assaying  2.85 
per  cent.,  or  14,152  pounds  fine  copper.  During  an  exposure  of 
186  days  the  ore  had  lost  34.3  per  cent,  of  its  copper. 

"All  the  experiments  made  on  a  total  of  nearly  3,000  tons  of 
ore  proved,  beyond  possibility  of  doubt,  an  average  loss  of  more 
than  one  unit  of  copper,  or  over  20  pounds  of  ingot  per  ton  of  ore. 
Tliis  great  loss  during  the  roasting  readily  accounted  for  the  deficit 
in  the  copper  production,  if  only  1^  percent,  was  deducted  from 
the  assay  value  of  the  ores  for  losses  by  treatment.  The  actual 
loss  by  the  smelting  process,  as  practised  at  Ducktown,  approached 
two  units.  Further  experiments  were  made  to  confirm  the  results 
obtained.  Experiments  in  roasting  in  furnaces  proved  that  no 
copper  escaped  in  the  fumes.  This,  indeed,  was  anticipated,  as 
the  heat  in  roasting  never  could  reach  a  point  at  which  copper  is 
volatile.  The  only  other  possible  loss  is  by  the  leaching  of  the 
roast-piles  during  the  heavy  rains  frequent  in  the  Ducktown  hills;, 
and  to  this  cause  the  great  losses  were  finally  ascribed.  In  refer- 
ring to  experiments  in  the  leaching  of  these  ores  later  on,  this 
subject  will  be  discussed  in  detail.  .Suffice  it  here  to  say,  that 
Avith  a  roasting  in  one  fire  only,  from  1  to  1^  units  of  copper  be- 
came soluble  in  water.  The  results  were  further  confirmed  by 
copper  found  in  large  quantity  in  the  clay  'bottoms'  of  the  roast- 
piles.  After  a  ^hower  of  rain,  the  roast-yard  would  be  covered 
with  pools  of  green  water  highly  charged  with  copper." 

During  the  last  two-thirds  of  the  life  of  the  roast-heap  it  hardly 
requires  an  hour's  labor,  and  if  the  works  possess  an  ample  stock  of 
roasted  ore  in  advance,  nothing  further  need  be  done  to  the  pile 
until  it  has  burned  itself  out  and  becomes  suffir^iently  cool  to  han- 
dle. The  daily  inspection,  however,  should  never  be  omitted;  for, 
even  at  this  advanced  stage  of  the  process,  irregular  settling  or  swell- 
ing of  some  portion  of  the  structure  may  cause  sufficient  Assuring 
and  consequent  admission  of  air  to  produce  serious  matting,  a  disas- 


ter  that  the  applicatioD  of  a  single  shovelful  of  fines  at  the  begiuuiug 
of  the  trouble  would  have  prevented.  lu  fact,  it  is  far  better  to 
leave  the  heap  undisturbed,  unless  good  reasons  exist  for  breaking 
into  it,  as  the  agglutinated  covering  material  forms  a  roof  almost 
impermeable  to  rain  and  wind,  while  the  freshly  calcined  ore, 
when  exposed  to  these  elements,  necessarily  undergoes  a  serious 
waste.  But  if,  as  is  in  most  instances  the  case,  the  demand  for 
ore  from  the  smelting  department  exceeds  the  supply  from  the 
mine,  but  scant  time  can  be  afEorded  to  the  intermediate  steps, 
and  the  calcination  must  suffer.  If,  therefore,  it  is  the  object  to 
utilize,  at  the  earliest  possible  moment,  the  ore  that  is  stored  up 
in  the  heaps,  they  should  be  closely  watched,  and  whatever  por- 
tions of  the  same — usually  the  ends  and  corners — are  found  to  be 
moderately  cool,  should  be  carefully  stripped  and  broken  into,  the 
object  being  to  cool  the  ore  that  is  already  roasted,  and  extinguish 
the  last  remains  of  fire  as  rapidly  as  possible,  without  interfering 
too  seriously  with  the  process  of  oxidation  that  is  continuing  in 
the  main  body  of  the  pile.  This  is  accomplished  by  digging  away 
the  calcined  o.?,  and  following  up  the  line  of  fire  as  it  recedes 
from  the  surface  toward  the  center,  without  approaching  it  so 
closely  as  to  completely  extinguish  it  in  that  portion  of  the  ore  not 
yet  properly  calcined,  which  is  easily  done  at  this  stage  of  the 
operation.  At  least  12  inches  should  be  left  between  the  outer  air 
and  the  line  of  active  oxidation,  and  it  is  a  good  practical  rule 
never  to  allow  the  surface  to  become  so  hot  as  to  be  unbearable  to 
the  naked  hand. 

The  too  common  practice  of  keeping  the  smelting  department 
so  far  in  advance  of  the  ore  supply  as  to  require  the  breaking  into 
and  utilization  of  roast-heaps  in  which  the  ore  is  still  red-hot,  and 
just  at  the  most  active  and  profitable  stage  of  calcination,  necessi- 
tates the  employment  of  a  strong  body  of  laborers  to  bring  water 
and  constantly  drench  the  smoking  ore,  in  order  to  make  it  at  all 
possible  for  the  other  workmen  to  shovel  it  into  their  barrows,  and 
must  be  condemned  as  unnecessary  and  productive  of  more  trouble 
and  expense  than  almost  any  other  practice  at  our  smelting  works. 

Among  these  sources  of  extra  expense  are  the  doubled  cost  of 
taking  down  and  transporting  the  roasted  material;  the  burning 
and  rapid  destruction  of  tools  and  cars;  the  medical  bills  claimed 
by  the  workmen  who  suffer  from  such  unhealthy  employment; 
.and,  far  greater  than  all.  the  injurious  effect  on  all  subsequent 

THE    ROASTING    OF    ORES    IN    LUMP   FORM.  127 

steps  of  the  process,  which  will  be  referred  to  in  the  chapter  cii 
"Smelting  in  Blast-Furuaces." 

On  the  other  hand,  the  only  possible  advantage  that  can  be 
claimed  is,  that  some  two  or  three  weeks'  interest  on  the  value  of 
the  ore  is  saved. 

When  the  heap  is  properly  cooled,  the  mass  of  ore,  which,  while 
still  hot,  is  often  almost  as  hard  and  tough  as  a  wall  of  solid  rock, 
crumbles  to  pieces  with  a  single  blow  of  the  pick,  ard  is  wheeled 
in  barrows  from  the  roast-heap  to  the  furnace  car. 

When  the  heap  is  sufficiently  cooled,  it  is  "'stripped"  by  remov- 
ing not  only  the  fines  that  formed  its  cover,  but  its  entire  surface, 
to  such  a  depth  as  is  necessary  to  include  all  material  that  has 
escaped  oxidation.  This  unroasted  material  is  made  up  largely  of 
the  fines  forming  the  cover,  and  which,  though  often  quite  thor- 
oughly oxidized  on  the  top  of  the  pile,  are  so  agglutinated  with 
sulphur  as  to  be  unfit  for  the  furnace.  The  covering  of  the  sides 
is  seldom  sufficiently  roasted,  and  this  is  especially  the  case  near 
the  ground,  where  the  ragging  itself,  to  a  depth  of  several  inches, 
is  frequently  found  unscathed.  The  angles  of  the  pile  are  also 
seldom  in  good  condition,  and  many  isolated  patches  and  bunches 
of  ore  will  be  found  that  the  careful  foreman  will  reject.  This 
statement,  however,  refers  rather  to  the  results  of  the  ordinary 
practice  than  to  those  that  can  easily  be  obtained  by  close  atten- 
tion to  details  and  by  enlisting  the  interest  of  some  intelligent 
foreman.  As  already  explained,  the  fire  will  find  its  way  to  every 
nook  and  corner  where  sulphides  still  exist,  if  only  the  conditions 
are  favorable.  The  author  recollects  with  satisfaction  the  morti- 
fication displayed  by  his  roasting  foreman  but  a  few  years  ago,  at 
the  unusual  occurrence  of  a  few  hundred-weight  of  fused,  and  a 
still  smaller  amount  of  raw,  ore  in  a  heap  of  some  200  tons. 

A  half  fused,  honeycombed  condition  of  the  upper  part  of  the 
heap,  presenting  the  appearance  of  a  skeleton  of  gangue  from 
which  all  mineral  has  been  melted  out,  is  a  certain  indication  of  a 
proportional  amount  of  matte  below.  This  molten  material  nat- 
urally gravitates  to  the  bottom  of  the  heap,  and  is  there  found  in 
masses  of  greater  or  less  extent;  often  of  many  tons'  weight, 
though,  in  such  a  case,  warning  would  have  been  given  during 
the  roasting  by  the  irregular  sinking  of  the  heap,  and  even  by 
depressions  and  crater-like  cavities  on  the  surface.  This  molten 
product  is  very  properly  termed  "  heap-matte,"  and  varies  but  little 
in  appearance  or  composition  from  the  similar  product  of  a  blast- 


furnace.  A  popular  impression  prevails  among  certain  foremen, 
and  even  assayers,  that  the  light  honeycombed  material  that  re- 
mains after  the  melting  ont  of  its  sulphide  constituents  is  rich  in 
copper,  but  the  contrary  is  true.  The  uufused  skeleton  merely 
represents  the  siliceous  shig,  while  the  molten  sulphide  mass  below 
is  the  equivalent  of  the  matte,  the  purity  and  value  of  either 
product  depending  on  the  temperature  to  which  the  ore  has  been 
subjected,  and  the  consequent  perfection  of  the  smelting  or  liqua- 
tion process.  This  fact  is  sustained  by  the  following  assays  of 
samples  of  considerable  size: 

No.  1.  No.  2. 

Original  ore  before  roasting 21.6  copper.  18.6  copper. 

Siliceous  skeleton T.3       "  6.4 

Heap-matte .   34.7       "  36  6        " 

The  formation  of  this  heap-matte  in  any  considerable  quantity 
is  very  detrimental  to  the  roasting  process,  but  is  easily  avoidable; 
for  it  is  invariably  caused  by  either  too  much  or  too  little  air.  In 
too  many  instances,  no  particular  notice  is  taken  of  its  occurrence, 
and  it  is  sent  to  the  smelting-furnace  mixed  with  the  well-roasted 
ore.  This  is  exceedingly  bad  practice,  and  should  on  no  account 
be  permitted,  as  it  is  totally  impossible  to  foresee  the  grade  of 
matte  that  will  be  produced  by  the  smelting  process  when  this 
nnroasfed  sulphide  is  mixed  in  unknown  and  varying  quantities 
with  the  properly  prepared  charge.  If  the  percentage  of  the  fur- 
nace mixture  be  such  tliat  the  addition  of  this  raw  matte  does  not 
lower  the  tenor  of  the  product  below  the  desired  standard,  it  may 
then,  of  course,  be  fed  with  the  roasted  ore,  but  should  be  kept 
strictly  by  itself,  and  added  to  each  charge  in  weighed  quantities. 
Any  infringement  of  this  i\ile  gives  rise  to  the  formation  of  a 
matte  varying  greatly  in  its  percentage  of  copper  as  well  as  in  its 
entire  composition,  and  deranges  not  only  the  smelting  process, 
but  seriously  affects  the  regularity  of  the  matte  concentration 

The  heap-matte  may  occur  in  such  masses  that  serioua  difficulty 
is  experienced  in  breaking  it  up,  especially  as  it  retains  its  heat  for 
a  great  length  of  time,  and  in  this  condition  is  almost  malleable, 
yielding  and  flattening  under  the  blows  of  the  sledge  like  a  block 
of  wronght-iron.  Much  expense  and  annoyance  mav  be  spared  by 
stripping  the  central  molten  mass  thoroughly  of  all  adhering  ore, 
and  allowing  it  to  cool  for  two  or  three  days;  at  the  expiration  of 
■which  time  it  will  be  found  quite  brittle  and  comparatively  easy  to 

THE    BOASTING    OF   OHES    IX    LUMP    FOliM.  129 

deal  with.  Thorough  and  repeated  drenchings  with  water  will 
produce  even  better  results;  but  it  should  be  borne  in  mind  that 
d  considerable  proportiou  of  the  cupriferous  contents  of  calcined 
ore  is  in  a  soluble  condition. 

When  through  carelessness  or  inexperience  heap-matte  is  formed, 
it  must  be  either  treated  together  with  the  matte  produced  from 
the  first  fusion  in  the  blast-furnace,  or  set  aside  until  a  sufficient 
amount  is  collected  to  form  a  small  heap  by  itself,  and  be  re- 
roasted.  It  should,  on  no  account,  be  mixed  with  the  raw  ore,  as 
it  demands  a  different  treatment,  and  will  either  cause  irregulari- 
ties in  the  ore-roasting,  or  will  pass  through  that  process  unaltered 
and  with  no  perceptible  diminution  in  its  percentage  of  sulphur. 

The  proportion  of  strippiugs  and  other  unfinished  products  of 
heap-roasting  that  may  be  considered  allowable  was  determined 
experimentally  by  simply  weighing  the  finished  and  unfinished 
portions  of  half  a  dozen  consecutive  roast-heaps,  averaging  about 
240  tons  each.  About  10  per  cent,  of  fines  were  used  for  the 
covering  layer  in  each  case.  The  total  amount  of  unroasted  mate- 
rial, as  given  in  the  following  table,  shows  that  even  a  portion  of 
the  fines  is  thoroughly  oxidized: 

Unroasted.  Roasted.        Days  Heap  was 

Per  Cent.  Per  Cent.  Active. 

No.  1...    9.6  90.4  64 

"     2 6.6  93.4  71 

"     3 8.4  91.6  70 

"     4 9.0  91.0  61 

"     5  7.6  92.4  67 

"     6 11.4  88.6  57 

The  figures  have  been  slightly  corrected,  without  altering  their 
relative  values,  to  make  the  aggregate  in  eacli  case  exactly  equal 
100  per  cent.,  which,  of  course,  can  never  be  precisely  attained  by 
addiug  the  weights  as  actually  arrived  at. 

While  these  results  are  taken  from  ordinary  everyday  work,  it 
should  be  understood  that  they  can  only  be  attained  by  the  most 
careful  attention  in  the  roasting-yard.  The  proportion  of  the 
product  rejected  as  unfit  for  the  smelting-furnace  at  some  works 
might  be  even  less  than  in  the  case  just  cited,  and  the  reason  may 
be  readily  recognized  in  the  low  grade  of  the  product  from  the 
fusion,  and  the  constant  complaints  of  the  impossibility  of  keep- 
ing the  matte  up  to  the  proper  standard.  A  selection  in  such 
cases  as  rigid  and  thorough  as  in  those  just  tabulated  would  result 
iu  the  rejection  of  from   25  to  60  jier  cent,  of  the  entire  heap. 


An  allowance  of  10  per  cent,  may  tlierefore  be  considered  reasonable 
— although  deniaudiog  more  than  ordinary  care  and  skill — and  of 
this,  three-fourths  sliould  be  tines.  Tlie  stripping  should  be  per- 
formed in  a  cleanly  and  systematic  manner,  and  to  an  extent  sev- 
eral feet  in  advance  of  the  line  of  excavation,  and  the  material  thus 
removed  piled  on  one  side,  to  be  subsequently  screened  on  the  first 
calm  day;  for  the  least  wind  causes  a  heavy  loss  when  handling 
this  half-oxidized  powder.  The  fine  part  is  again  used  as  a  cover- 
ing, for  which  it  is  much  better  suited  than  raw  ore,  while  the 
much  smaller  coarse  portion  is  added  to  the  nearest  heap  in  process 
of  erection. 

It  will  be  readily  seen  that  very  much  more  fine  ore  is  produced 
during  the  processes  of  mining  and  crushing  than  can  be  used  for 
the  purpose  of  covering  material,  especially  as  only  a  small  pro- 
portion of  the  latter  is  sufficiently  oxidized  at  each  operation  to 
be  passed  on  to  the  smeltiug-furnace.  The  problem  of  the  best 
means  of  utilizing  this  constantly  increasing  amount  of  fine  ore  in 
works  unprovided  with  calcining-furnaces  is  often  a  pressing  one. 
It  will  be  referred  to  again,  under  the  heading,  "The  Treatment 
of  Pulverized  Ores." 

The  roast-heap,  when  once  tolerably  cool,  is  torn  down  and 
loaded  'nto  the  furnace-car  with  great  celerity.  Three  or  four 
men  trundle  the  barrows,  while  double  that  number  wield  the 
pick,  shovel,  and  hammer.  It  is  the  duty  of  these  laborers  to 
break  all  partially  fused  masses,  or  lumps  that  are  too  large  for 
proper  smelting,  into  fragments  of  a  reasonable  size,  as  especially 
determined  by  the  metallurgist.  There  is  not  time,  or  space,  or 
opportunity  on  the  charging  floor  of  a  blast-furnace  in  full  opera- 
tion to  attend  to  any  duties  beyond  those  immediately  connected 
with  weighing  the  charge  and  filling  the  furnace,  and  many  serious 
irregularities  in  the  smelting  may  be  traced  to  an  omission  of  this 
simple  and  obvious  precaution. 

A  careful  and  humane  foreman  can  do  much  to  mitigate  the 
annoy.jnce  and  suffering  to  which  the  workmen  are  subjected 
during  the  labor  of  tearing  down  a  heap,  by  moving  the  point  of 
attack  from  one  to  the  other  side  of  the  pile,  according  to  the 
direction  of  the  wind,  as  well  as  by  keeping  the  fresh  surface  on 
which  the  men  are  engaged  well  sprinkled  with  water  to  settle  the 
fine  ore-dust.  At  best,  this  labor  is  the  most  disagreeable  and 
wearing  connected  with  ordinary  smelting,  and,  if  possible,  laborers 
should  be  changed  periodically  to  some  other  employment. 

THE    ROASTING    OF    ORES    IN^    LUMP    FORM.  131 

Aside  from  the  coiumou  tools  already  enumerated,  long,  stout 
steel  gads  and  a  few  heavy  sledges  are  needed  to  break  up  the 
central  portion  of  the  structure,  which,  although  not  fairly  fused, 
is  often  so  stuck  together  as  to  require  considerable  labor  for  its 
removal.  At  no  other  work  are  shovels  so  rapidly  destroyed,  and 
it  is  to  this  place  that  all  partially  worn,  though  still  serviceable, 
tools  are  sent  to  terminate  their  existence. 

The  tearing  down  of  the  heap,  and  breaking-up  of  the  matte 
that  may  be  formed  in  it,  are  greatly  facilitated  by  the  use  of  a 
small  quantity  of  dynamite,  or  other  high  explosives,  selecting  a 
powder  of  rather  low  force;  containing  not  over  30  per  cent,  of 

When  this  is  properly  used  and  in  not  too  hirge  quantities,  it 
saves  infinite  labor  with  bar  and  pick,  a  single  shot,  placed  in  a 
hole  made  in  half  a  moment's  time  with  a  bar,  often  accomplish- 
ing more  than  hours  of  hard  labor.  The  shot  should  sim.ply  shake 
up  and  loosen  the  mass,  leaving  the  large  lumps  to  be  broken  up 
by  sledge  and  pick,  as  usual.  If  enough  powder  is  used  to  break 
the  whole  mass  up  into  small  fragments,  a  great  portion  of  the  ore 
will  soar  into  the  air  and  go  toward  top-dressing  the  surrounding 

I  have  never  been  able  to  get  my  men  to  be  economical  enough 
with  their  powder,  except  by  forcing  them  to  pay  for  it  them- 
selves. When  they  realize  that  every  penny  that  is  saved  on  pow- 
der goes  into  their  own  pockets,  it  is  astonishing  how  little  it 
takes  to  do  the  same  work  that  required  several  times  the  quantity 
when  it  cost  them  nothing. 

After  the  complete  removal  of  the  old  heap,  and  any  slight 
repairs  that  may  be  required  to  restore  the  ground  to  its  former 
level,  a  thin  layer  of  raw  fines  is  again  spread  on  the  old  spot,  and 
the  fuel  arranged  for  a  fresh  pile.  The  estimate  of  costs  for  this 
process,  as  given  below,  is  based  on  many  different  ores,  varying 
greatly  in  composition,  and  under  very  various  circumstances,  and 
is  purposely  made  somewhat  liberal  to  allow  for  the  occurrence  of 
bad  work  and  various  other  mishaps  that  are  certain  to  occur  in  a 
greater  or  less  degree.  It  is  based  upon  a  plant  of  200  tons  daily 
capacity,  and  on  the  assumption  of  only  a  short  distance  for  trans- 
portation of  the  roasted  ne  to  the  smel ting-furnace.* 

♦This  is  a  considerable  reduction  on  tbe  original  estimates  for  this  process, 
s  published  in  the  previous  edition  of  this  work. 


Estimate  for  Roasting  200  Tons  Ore  per  24  Hours. 

Transportation  by  gravity-road  at  4i  cents  per  ton $9.00 

Labor  in  building  and  burning  heaps: 

6  men  at  |l.oO  =  $9.00 

2  men  at  $2.00  =  $4.00 


Five  cords  (640  cubic  feet)  wocid  at  $5.00 25.00 

Removing  and  loading  roasted  ore  by  contract  at  12  cents  per  ton 24  OO 

One  foreman 2. 50 

Screening,  patciiing  yard,  etc.,  2  men  at  $1.50   3  OO 

Oil,  lights,  repairs  to  cars,  track,  and  tools,  and  new  tools 11.50 

Transportation  to  furnace  in  dump  cars 9.00 

Total $97.00 

Or  $0.48i  per  ton  raw  ore. 

The  various  operatious  of  heap-ro:isting  may  ofteu  be  performed 
by  contract  to  great  advantage,  especially  if  one  has  a  good  fore- 
man to  see  that  the  quality  of  the  roast  is  kept  up  to  a  satisfactory 

To  give  an  idea  of  the  prices  that  are  fair  for  this  operation,  I 
will  mention  what  I  paid  for  roasting  a  heavy,  pyrrliotite  ore  in 
large  quantities,  say  150  to  "200  tons  per  day,  the  climate  being 
excessively  cold  and  stormy,  and  laborers'  wages  about  ^1.40  per 
10  hours.  The  company  furnished  the  wood  for  the  roast-beds, 
and  delivered  the  cars  at  the  yard;  but  the  cars  had  to  be  unloaded 
by  hand  and  the  raw  ore  wheeled  to  the  heaps,  the  arrangements 
for  dumping  the  ore  direct  not  having  been  then  completed. 

For  unloading  the  raw  ore  on  the  heaps,  laying  the  wood,  com- 
pleting heaps,  and  covering  and  watching  them  throughout  the 
entire  operation,  $0.22  per  ton  of  ore. 

For  stripping,  tearing  down,  and  loading  the  roasted  ore  on 
cars,  and  unloading  the  cars  by  hand  into  the  smelter-bins,  $0.10 
per  ton  of  ore. 

The  contractors  furnished  their  own  powder,  but  the  company 
provided  tools,  barrows,  etc.,  though  the  contractors  paid  for  the 
sharpening  of  their  bars,  picks,  etc. 

On  the  above  basis,  the  oontractors  made  a  fair  profit  when  they 
attended  strictly  to  their  business,  and  when  there  were  no  inter- 
ruptions or  shiit-dowu.  The  degree  of  desiilphurization  arrived  at 
by  this  process  is  seldom  accurately  determined,  owing  to  the  ditfi- 
cnlty  and  expense  of  obtaining  an  accurate  sample,  and  to  the  fact 
that  the  experienced  eye  can  very  correctly  judge  of  the  success  of 

THE    ROASTING    OF    ORES    IN    LUMP    FORM.  133 

tlie  roast,  while  auy  defect  in  the  ijrocess  will  become  immediate]} 
apparent  iu  the  lower  tenor  of  the  product  of  the  succeeding  fusion. 
Owing  to  the  scarcity  of  accurate  investigations  on  the  subject,  tho 
following  determinations  were  made: 

No.  1.  A  heavy  pyritous  ore,  from  the  Ely  mine,  Vermont, 
consisting  principally  of  magnetic  pyrites  and  chalcopyrite,  burned 
in  a  heap  of  about  300  tons  for  eleven  weeks.  After  stripping  off 
the  surface,  a  sample  of  the  roasted  ore,  as  delivered  at  the  smelt- 
ing-furuac€,  was  taken.  The  following  was  the  assay  of  the  ore 
before  and  after  calcination : 

Before  Roasting.  After  Roasting. 

Sulphur 33.6  per  cent.  7.4  per  cent. 

Copper 8.2        "  9.1 

Insoluble 27.0         "  31.1 

The  condition  of  the  copper  in  the  roasted  sample  was  also 
■determined  in  this  case,  as  follows: 

Sulphate   of   copper 1.3  per  cent. 

Oxide  of  copper 2.1         " 

Sulphide  of  copper 5.7         " 

Total 9.1 

No.  2.  A  heavy  pyritous  ore,  being  almost  pure  iron  pyrites 
containing  minute  quantities  of  copper,  silver,  and  gold,  from  the 
Phillips  mine.  Buckskin,  Colorado,  was  roasted  for  6  weeks  in 
piles  of  60  tons,  and  was  used  as  a  flux  for  siliceous  silver  ores.  A 
careful  sample  of  the  roast  yielded  sulphur,  before  roasting,  -iG^ 
per  cent. ;  after  roasting,  11  per  cent. 

A  considerable  number  of  similar  tests  give  corresponding  re- 
sults, showing  that  a  very  fair  degree  of  desulphurization  can  be 
attained  by  this  crude  and  ancient  method,  but  still  better  results 
will  be  reached  in  ores  containing  less  pyrites,  and  making  the' 
fact  evident  that,  in  heap-roasting  as  well  as  in  the  calcination  of 
pulverized  sulphides,  the  copper  is  the  last  metal  present  to  part 
with  its  sulphur,  and  that  a  large  proportion  of  this  still  remains 
ill  the  condition  of  a  sulphide  after  nearly  the  entire  iron  contents 
have  become  thoroughly  oxidized.  This  agrees  perfectly  with  all 
investigatioES  relative  to  the  comparative  affinity  of  sulphur  for 
the  various  metals,  and  is  iu  no  other  motallurgical  process  more 
•strikingly  exemplified  than  in  the  so-called  "  kernel-roasting,"  as 
practised  at  Agordo,  iu  Italy.     There,  the  mechanical  separation 


of  the  copper  from  its  accompauyiug  pyritous  gangue  is  effected 
by  stopping  the  process  of  calciuatioii  at  the  exact  point  where  the 
entire  iron  contents  have  been  oxidized  into  a  soft  earthy  material, 
while  the  copper  remains  in  combination  with  snlphur  in  a  hard, 
metallic  condition,  and,  most  singularly,  retreats  into  the  center 
of  each  lump  of  ore,  forming  a  heavy  and  solid  kernel,  which  can 
easily  be  separated  from  its  earthy  envelope  by  inexpensive  me- 
chanical means.  As  this  interesting  process  is  not  practised  in 
this  country,  and  in  all  probability  is  not  suited  to  our  domestic 
conditions,  the  student  desirous  of  pursuing  the  subject  will  find 
further  information  in  Plattner's  liostjirocesse,  as  well  as  in  a 
paper  by  the  author  in  the  Mineral  Resources  of  the  Cnited  States 
(A.  Williams,  Jr.,  1883). 

During  the  past  few  years,  very  much  better  results  have  been 
obtained  in  heap-roasting  than  would  have  formerly  been  consid- 
ered possible.  Ores  containing  over  40  per  cent,  sulphur  are  now 
often  roasted  down  to  7  or  8  per  cent.,  with  regularity  and  cer- 
tainty. This  comes  partly  from  longer  experience  of  svorkmen, 
partly  from  premiums  paid  the  men,  based  on  the  grade  of  matte 
produced  in  the  subsequent  smelting  operation,  and  partly  from 
using  a  much  less  quantity  of  wood  to  kindle  the  heap,  and  con- 
ducting the  entire  operation  of  roasting  in  a  much  more  repressed 
and  gradual  manner. 

The  appearance  of  a  freshly-opened  heap  of  well-roasted  ore  is 
characteristic,  athough  difficult  of  description.  It  should  present 
a  strictly  earthy,  irregular  surface  of  a  blackish-brown  hue,  the 
scarcity  of  air  preventing  the  oxidation  of  the  iron  to  the  red  ses- 
quioxide.  This  is  a  decided  advantage  in  a  reverberatory  smelting- 
furnace,  where  the  powerful  carbonic  oxide  atmosphere  of  the 
blast-furnace  is  wanting  to  reduce  it  to  the  protoxide  and  thus 
fit  it  for  entering  the  slag,  the  higher  oxide  being  infusible  at  ordi- 
nary smelting  temperatures.  It  is,  in  fact,  principally  a  magnetic 
oxide,  and,  while  the  greater  part  of  the  contents  should  adhere 
closely  together,  and,  when  disturbed,  should  come  out  in  the 
shape  of  large  lumps,  no  sign  of  actual  fusion  should  be  visible, 
and  the  largest  mass  should  fall  into  fragments  at  a  few  blows  of 
the  hammer.  The  more  siliceous  pieces  of  ore  will  have  taken  on 
a  somewhat  milky  and  opaque  look  in  place  of  the  ordinary  vitre- 
ous appearance  of  quartzose  minerals,  and  the  veinlets  of  sulphides 
traversing  the  same  will  be  found  oxidized  throughout.  The  solid 
lumpi  of  pyrites,  if  carefully  broken,  will   usually  display  a  series 

THE    ROASTING    OF   ORES    IX    LUMP   FORM.  135 

of  coDcentiic  layers,  completely  oxidized  aud  earthy  on  the  out- 
side, aud  gradually  acquiring  greater  firmness  and  a  slight  sub- 
metallic  luster,  which  culminates  in  a  rich  kernel  near  the  center 
of  the  fragment.  This  resembles  strongly  one  or  other  of  the 
grades  of  matte  as  produced  from  the  smelting-furnace,  and  usu- 
ally contains  the  greater  part  of  the  entire  copper  contents  of  the 
lump.  The  silver — if  any  be  present — is  also  concentrated  in  a 
marked  degree,  though,  so  far  as  the  author's  own  investigations 
extend,  not  with  the  same  remarkable  perfection  as  the  less  pre- 
cious metal.  The  examination  of  a  characteristic  lump,  such  as 
just  described,  which  contained  before  roasting  about  4  per  cent, 
of  copper,  yielded  the  following  interesting  results: 

The  outer  eartlily  envelope  contained. . . .  Traces  of  copper. 

The  medium  concentric  layers 1.2  per  cent.         " 

The  central  sub-metallic  kernel 69.6         "  " 

An  imperfect  roasting  is  quickly  detected  by  the  presence  of 
more  or  less  fused  material  at  certain  portions  of  the  heap,  while 
elsewhere  there  exists  no  cohesion  between  the  lumps  of  ore,  which 
fall  a])art  like  so  many  paving-stones.  A  certain  metallic  appear- 
ance will  also  be  noticed,  very  different  from  the  dull,  earthy  char- 
acter of  the  projierly  burned  pile.  Although  a  large  proportion  of 
the  contents  may  exhibit  quite  a  brilliant  red  color,  as  though  an 
unusually  perfect  oxidation  of  the  iron  had  taken  ])lace,  a  mere 
weighing  of  one  of  the  lumps  in  the  hand  will  quickly  undeceive 
the  least  experienced  observer,  and  its  fracture  will  show  that  the 
effect  of  the  fire  was  only  surface  deep,  while  the  entire  interior 
remains  unaltered.  A  careful  study  of  diflferent  roast-heaps, 
wherever  opportunity  offers,  will  soon  render  the  student  skillful 
in  judging  by  eye  of  the  degree  of  success  attained  by  tliis  process, 
and  in  after-life  frequently  furnish  him  the  key  to  the  cause  of 
the  unsatisfactory  tenor  of  the  matte  produced  from  his  furnacei=. 
No  metallurgical  process  is  more  dependent  upon  an  efficient  and 
conscientious  foreman,  and  the  best  results  are  usually  obtained 
by  selecting  some  intelligent  and  ambitious  man  from  the  roast- 
yard  laborers,  and  holding  him  strictly  responsible  for  results. 

A  decided  improvement  in  heap-roasting  of  ores  was  introduced 
at  the  works  of  The  Canadian  Copper  Com]iany  of  Sudburv, 
Ontario,  under  the  managotnent  of  the  author,  in  1888-89.  It 
was  first  tried  by  his  assistant,  Mr.  James  McArthur,  and  proved 


SO  valuable  that  it  became  a  regular  practice  under  ordinary 

We  have  called  it  the  "V-Method"  of  roasting,  and  the  accom- 
panying sketch  will  make  it  clear.  It  consists  in  introducing  a 
supplementary  roast-heap  between  each  two  regular  heaps,  so  that, 
if  left  untouched,  there  would  he  a  continuous  and  unbroken 
roast-heap  the  entire  length  of  the  roast-yard. 

The  supplementary  heap  should  not  be  built  until  its  two  neigh- 
bors, which  are  to  form  its  lateral  walls,  are  well  under  way,  and 
have  been  lighted  from  10  to  14  days.  By  this  time,  if  properly 
managed,  they  will  be  cool  enough  on  the  outside  to  run  no  risk 
of  setting  afire  the  bed  of  wood  which  is  laid  down  for  the  supple- 
mentary heap.  The  fresh  bed  of  wood  is  laid  down  much  thinner 
than  for  independent  heaps,  and  a  single  layer  is  extended  well 
up  the  slope  of  the  two  neighboring  heaps.  The  ore  is  dumped 
on  as  rapidly  as  possible,  and  the  heap  finished  off  with  ragging 
and  fines  in  the  usual  manner,  and  fired  from  the  ends. 

No.  I.  No.  3.  No.  2. 

Fig.  19. — The  Y-Method  of  Heap-Roasting. 

The  result  is  excellent,  for  the  new  heap,  having  its  sides  pro- 
tected, burns  clear  through  its  entire  extent,  and  then  sets  on  fire 
the  still  unroasted  ore  on  the  outside  of  the  two  neighboring  heaps. 

Thus  the  proportion  of  unroasted  ore  is  reduced  to  a  minimum, 
and  indeed  is  seldom  worth  keeping  separate. 

Another  great  advantage  is  the  economizing  of  space,  for  by  this 
arrangement  some  60  per  cent,  is  added  to  the  capacity  of  the 

It  may  require  some  little  patience  and  experimentation  at  first 
to  adapt  this  practice  to  a  new  ore,  but  it  is  well  worth  the  trouble, 
and  has  been  pronounced  by  various  members  of  our  profession  a 
decided  and  important  improvement  in  this  ancient  and  useful 

In  the  case  referred  to,  the  ore  that  was  roasted  was  a  nickelif- 
erous  pyrrhotite  mixed  with  chalcopyrite;  but  I  have  tried  it  suffi- 
ciently on  both  heavy  and  lean  ores  of  the  ordinary  yellow  iron 
pyrites  to  know  that  it  is  equally  well  adapted  to  all  ores  that  are 
any  way  suited  to  heap-roasting. 

THE    KOASTIXG    OF    ORES    IN    LUMP   FORM.  137 


There  remains  only,  in  connection  vvitli  this  portion  of  the  sub- 
^"ect,  to  notice  the  slight  deviations  that  it  is  found  necessary  to 
introduce  in  adapting  this  method  to  the  treatment  of  mattes. 

These  artilicially  formed  sulphides,  containing  variable  percent- 
ages of  sulphur,  may  be  sufficiently  desulphurized  in  heaps,  and 
their  chemical  composition  has  no  marked  effect  upon  the  result, 
25rovided  lead  is  not  present  to  such  an  extent — 15  per  cent,  or 
more — as  to  increase  the  fusibility  of  the  material. 

The  most  marked  distinction  between  the  behavior  of  ore  and 
matte,  when  submitted  to  this  process,  is  the  fact  that,  while  the 
former  substance  may  be  satisfactorily  oxidized  by  a  single  treat- 
ment, the  latter  invariably  demands  two,  and  oftener  three  or 
more  separate  burnings,  before  it  is  properly  prepared  for  the  suc- 
ceeding fusion.  There  is  no  exception  to  this  rule,  wiiich,  if  prop- 
erly understood,  would  prevent  the  disappointment  frequently 
experienced  by  those  unaccustomed  to  this  method  of  desulpluuiz- 
ing  matte  and  who  are  led  to  condemn  the  practice  on  finding,  at 
the  conclusion  of  the  first  carefully  conducted  burning,  that  the 
•only  visible  results  are  a  slight  scorching  of  the  surface  of  each 
fragment,  a  change  in  color  from  the  original  brownish-black  to  a 
brassy  yellow,  and  a  more  or  less  extended  fusion  of  such  portions 
of  the  heap  as  have  sustained  the  greatest  heat.  In  reality,  the 
influence  of  the  process  has  been  much  more  profound  than  can  be 
realized  from  external  appearances,  and  although  neither  the  re- 
moval of  the  sulphur  nor  the  oxidation  of  the  iron  and  copper  has 
progressed  to  any  great  extent,  a  certain  change  in  the  physical 
condition  of  every  fragment  of  matte  has  been  effected  that  pre- 
pares it  perfectly  for  a  second  burning,  and  which  seems  a  neces- 
sary preliminary  to  the  actual  desulphurization. 

Each  succeeding  operation  requires  a  slightly  increased  propor- 
tion of  fuel,  as  the  volatilization  of  the  sulphur  and  the  oxidation 
of  the  metallic  constituents  deprive  the  matte  of  its  internal 
sources  of  heat,  and  at  the  same  time  greatly  lessen  its  fusibility. 

For  the  first  roasting,  a  bed  of  wood  should  be  prepared  similar 
to  that  for  a  heap  of  ore,  although  smaller  in  area;  for  it  is  diffi- 
cult to  regulate  the  temperature  and  prevent  matting  in  a  heap 
much  larger  than  12  feet  square,  and  this  will  be  found  a  conven- 
ient size  to  hold  from  60  to  70  tons  of  matte  when  raised  to  a 
height  of  about  6  feet.     A  single  chimney  in  the  center  is  suffi' 


cieut,  aud  about  this  structure  the  broken  matte  should  be  heapec) 
just  as  it  comes  from  the  crusher  or  spalliug-floor,  aud  regardless 
of  the  fines  that  it  contains.  The  presence  of  these  has  been 
fouud  necessary  to  check  the  rapidity  of  the  operation,  and  pre- 
vent the  fire  from  suddenly  spreading  through  the  entire  pile  in  a 
few  hours  without  accomplishing  any  useful  result,  though  gener- 
ating for  a  short  time  a  temperature  high  enough  to  fuse  a  large 
proportion  of  the  contents  into  a  single  lump. 

Less  care  need  be  taken  in  siiapiug  a  matte-heap  than  in  the 
case  of  ore,  and  it  is  merely  necessary  to  build  it  up  in  the  form  of 
a  rude  mound,  which  may  best  be  covered  with  thoroughly  burned 
ore  from  the  roast-heaps,  most  of  which  on  handling  will  crumble 
to  a  sutficieiit  fineness  for  the  purpose,  while  any  hard  lumps  may 
be  removed  with  the  dung-fork.  This  obviates  any  screening  or 
classifying  of  the  matte  in  the  open  air,  which  always  entails  a 
heavy  loss,  owing  to  the  great  value  aud  excessive  friability  and 
lightness  of  the  material  after  calcination.  If,  as  is  usually  the 
case,  the  proportion  of  fines  after  the  first  burning  is  found  so 
great  as  to  endanger  the  proper  combustion  of  the  heap  for  the 
second  operation,  the  mechanical  loss  may  be  reduced  to  a  mini- 
mum by  separating  the  excess  of  pulverized  matte  by  the  use  of  a 
dung-fork,  with  tines  closely  set,  during  the  turning  of  the  ore 
from  the  heap  just  finished  on  to  the  fresh  bed  of  wood,  and  at 
the  conclusion  of  the  process  removing  the  fines  that  are  thus 
isolated,  either  directly  to  the  snielting-house,  or,  if  they  still  con- 
tain too  much  sulphur,  to  the  calcining-furnaces.  The  covering 
of  the  original  heap,  consisting  solely  of  roasted  ore,  should  be 
stripped  off,  and  either  sent  to  the  smelting-furnace  or  again  used 
for  a  similar  purpose.  It  need  hardly  be  mentioned  that  the 
presence  of  arsenic  or  similar  impurities  in  the  ore,  in  greater 
quantities  than  in  the  matte,  should  prevent  any  "^uch  practice  as 
that  just  recommended,  and  it  may  be  accepted  as  a  universal  rule 
in  copper  smelting,  that  no  impure  ores  or  products  should  ever 
be  mixed  with  those  freer  from  deleterious  substances. 

Under  no  circumstances  need  a  matte-pile  be  covered  as  thor- 
oughly as  a  roast-heap  consisting  of  ore,  nor  can  the  formation  of 
a  considerable  amount  of  matte,  which  in  ore-roasting  would  be 
evidence  of  a  great  want  of  skill  or  care,  be  considered  as  a  re- 
proach, experience  having  so  conclusively  shown  the  impossibility  of 
preventing  its  occurrence  tliat,  unless  about  one-eighth  of  the  lower 
portion  of  a  matte-heap  isthusfused,  no  thorough  oxidation  of  the 


remamder  will  be  effected.  The  time  necessary  for  the  operations 
just  discussed  varies  according  to  the  quality  of  the  matte,  the 
condition  of  the  weather,  and  certain  other  factors,  but  will  in 
general  be,  for  tlie  first  burning,  eight  days,  while  on  the  tenth 
day  the  heap  will  be  sufficiently  cool  to  permit  its  turning  on  to  a 
fresh  layer  of  fuel.  The  second  operation  requires  a  day  longer, 
and  the  third  a  day  less  than  the  first  burning. 

To  those  familiar  with  the  practice  of  heap-roasting  as  applied 
to  ores,  no  particular  directions  are  necessary  except  that  care 
should  be  taken  that  the  large  blocks  of  matte  that  are  formed 
during  each  burning  be  well  broken  np  and  placed  near  the  center 
of  the  heap  next  constructed,  that  they  may  have  every  opportunity 
for  a  thorough  desulphurization. 

Whatever  raw  matte  still  remains  from  the  last  burning  is  best 
reserved  until  the  construction  of  a  fresh  heap  furnishes  the  proper 
means  for  its  treatment.  At  the  last  two  burnings,  it  is  well  to 
introduce  two  or  mure  layers  of  chips,  bark,  or  other  refuse  fuel, 
into  the  matte-heap;  for  it  will  act  powerfully  in  decomposing 
the  sulphates  that  at  this  stage  are  formed  in  considerable  amount, 
and  also  exercise  a  similar  and  most  marked  effect  on  whatever- 
compounds  of  arsenic  and  antimony  may  be  present.  This  simple 
measure  had  a  sufficient  effect  in  a  certain  instance  in  the  experi- 
ence of  the  author  to  be  plainly  noticeable  in  the  quality  of  the 
ingot  copper  produced. 

No  attempt  to  select  such  portions  of  thoroughly  calcined  mate- 
rial as  will  be  found  after  the  second  burning  has  ever  proved  remu- 
nerative. The  heap  of  matte  must  be  treated  as  a  whole,  and  the 
roastings  continued  until  the  desired  grade  of  desulphurization  is 

The  process  just  described  is  seldom  an  advantageous  one,  as^ 
aside  from  the  production  of  the  vilest  fumes  known  to  metallurgy, 
the  value  of  the  material  operated  on  is  too  great  to  admit  of  being 
locked  up  for  30  days  or  more,  or  to  warrant  the  loss  that  neces- 
sarily results  from  such  frequent  handling  in  the  open  air.  The 
hist  difficulty  may  be  partially  obviated  by  erecting  a  light  struc- 
ture to  protect  the  heaps  from  the  rain  and  wind;  but,  at  best, 
the  practice  is  an  imperfect  and  objectionable  one,  and  only  to  be- 
recommended  in  new,  outlying  districts,  where  an  expensive  cal- 
cining plant  cannot  at  once  he  erected,  and  where  the  climate  is 
favorable  for  out-of-door  operations.  The  expense  of  crushing 
and  calcining  in  furnaces  is  decidedlv  less  than  the  three  or  four 


t)iirnings  necessary  to  produce  the  same  result;  but  the  coDclition 
of  the  roasted  material  is  so  much  more  favorable  for  the  succeed- 
iug  smeltiug  process,  in  the  case  of  heap-roastiug,  that  this  reasou 
aloue  is  often  sufficieut  to  outweigh  all  objections  that  can  be 

The  practice  of  spalliug  the  large  pieces  of  matte  upon  the  heap 
itself  must  be  deprecated,  as  it  has  a  strong  tendency  to  solidify 
the  structure  and  render  the  draught  weak  and  irregular. 

The  cost  of  this  process,  based  upon  the  roasting  of  many  thou- 
sand tons  of  matte,  and  divested  of  those  details  that  too  closely 
resemble  the  heap-roasting  of  ore  to  warrant  repetition,  is  as  fol- 
lows, assuming  the  daily  amount  of  fresli  matte  subjected  to  this 
treatment  to  average  30  tons: 

COST   PER    TON    OF    MATTE. 

First  Fire. 

Breaking $0.19 

Transportation  to  lieap 0.06 

Fuel — allowing  3  cords  of  wood  to  60  tons  of  matte 0.25 

Constructing  heap  and  burning 0.21 

Total fO.71 

Second  Fire. 

Fuel — same  as  before  witli  addition  of  chips $0.30 

Turning  heap  and  burning ...     0.26 

Total 10.56 

Third  Fire. 

Fuel — same  as  second  fire |0.30 

Removing  finished  heap 0.23 

Transportation  to  furnace  and  expense  of  preparing  the  raw 
matte  still  remaining,  which  results  from  the  fused 
matte 0.26 

Total 10.78 

Total  cost  of  three  burnings $2.05 


At  jnst  what  period  in  the  history  of  the  art  it  became  cnstomary 
to  inclose  the  roast-heap  with  a  little  wall  of  earth  or  mason-work, 
in  order  to  protect  it  against  the  elements,  to  concentrate  the  heat, 
and  to  render  unnecessary  the  tedious  labor  of  covering  the  sides 
with  fine  ore,  is  unknown,  though  Agricola's  work  on  metallurgy 
shows  that  it  was  no  novelty  in  the  sixteenth  century.  These 
simple  walls  have  since  been  heightened  and  sometimes  connected 

THE    KOASTING    OF    ORES    IX    LUMP    FORM.  141 

with  au  arclKitl  roof;  the  area  that  they  inclose  has  been  jiaved 
and  occasionally  furnished  with  a  permanent  grate;  and,  more 
important  than  all,  the  interior  of  the  stall  has  been  connected  by 
a  flue  with  a  tall  chimney,  by  which  the  draught  has  been  im- 
proved, thus  shortening  the  process  of  oxidation,  while  the  noxious 
fumes  are  discharged  into  the  atmosphere  at  such  a  height  as  to 
render  them  unobjectionable  in  most  cases. 

A  very  great  variation  exists  in  the  size,  shape  and  general 
arrangement  of  stalls,  hardly  two  metallurgical  establishments 
building  them  after  the  same  pattern,  though  all  essential  differ- 
ences may  be  properly  considered  by  dividing  them  into  two 

1.  Open  stalls,  suitable  only  for  ore. 

2.  Covered  stalls,  suitable  for  both  ore  and  matte. 

1.  Open  Stalls. — Any  attempt  at  an  exhaustive  description  of 
the  diff'ereut  patterns  of  ore-stalls  that  human  ignorance,  as  well 
as  ingenuity,  has  invented,  would  be  a  waste  of  space.  They  all 
consist  of  a  comparatively  small  paved  area,  surrounded  by  at 
least  three  permanent  walls,  and  usually  having  an  open  front, 
which  is  loosely  built  up  at  each  operation,  to  confine  the  contents. 
The  back  or  sides,  or  both,  are  pierced  with  small  openings  com- 
municating with  a  flue  common  to  a  large  number  of  stalls  that 
enters  a  high  stack.  Tlie  draught  is  confined  to  these  passages  by 
covering  the  surface  of  the  ore  with  a  layer  of  fines.  From  the 
great  variety  of  existing  patterns,  one  built  at  the  works  of  the 
Parrot  Copper  and  Silver  Company,  of  Butte  City,  Montana,  is 
selected  for  description  as  possessing  exceptional  advantages  as  re- 
gards cheapness  of  construction,  convenience  of  filling  and  empty- 
ing, economy  of  fuel,  and  general  adajatability. 

The  stalls  may  be  built  either  of  common  red  brick,  of  stone, 
or,  far  better,  of  slag  molded  into  large  blocks,  which,  from  their 
size  and  weight,  require  little,  or  no  extraneous  support;  while 
brick  demands  thorough  and  extensive  tying  together  with  iron- 
work, and  stone  of  proper  size  and  shape  is  expensive  and  is  apt  to 
crack  when  exposed  to  great  fluctuations  of  temperature. 

As  these  so-called  "slag-bricks"  are  invaluable  for  walls  and 
foundations,  and,  in  fact,  for  every  purpose  for  which  the  most 
expensive  cut  granite  would  prove  available,  and  as  they  can  be 
produced  from  almost  any  copper  slag  that  is  not  too  basic,  a  brief 
description  of  the  cheapest  and  best  method  of  manufacturing 
them  is  appended. 



These  are  generally  made  from  the  slag  of  reverberatory  smelt- 
iug-furnaces,  both  because  this  material  is  usually  more  siliceous 
th;m  any  other,  and  also  because,  during  the  process  of  skimming, 
jt  can  be  obtained  in  large  quantities  in  a  very  brief  space  ot  time. 
There  should  be  no  difficulty,  however,  in  making  the  brick  froui 
the  slag  of  a  blast-furnace,  provided  the  smelting  is  sufficiently 
rapid  to  fill  the  molds  properly,  and  that  it  is  not  so  basic  as  lo 
yield  too  fragile  a  material  on  cooling.  Even  with  exceedingly 
brittle  blocks,  produced  from  a  highly  ferruginous  ore,  excellent 
and  durable  walls  can  be  constructed,  provided  the  blocks  are 
placed  in  position  uninjured;  for  they  will  bear  an  immense  crush- 
ing weight  with  impunity,  and  seem  to  defy  the  action  of  the 

Assuming  the  slag  to  be  obtained  from  a  reverberatory  furnace, 
the  process  of  preparing  the  molds  should  be  begun  as  soon  as  pos- 
sible after  the  slabs  from  the  previous  skimming  have  been  removed 
and  all  chips  and  fragments  cleared  from  the  sand  bed  by  the  aid 
of  a  close-toothed  iron  garden-rake.  Ordinary  loam — or  a  natural 
mixture  of  fine  sand  and  clay  of  such  consistence  that,  when 
slightly  moistened,  it  will  ball  firmly  in  the  hand — is  the  proper 
-material  for  the  molds,  which  should  be  formed  by  means  of  a 
number  of  wooden  blocks,  of  the  required  size,  carefully  smoothed 
and  slightly  tapered  to  facilitate  their  removal  from  the  sand,  and 
furnished  with  a  30-inch  handle,  inserted  in  their  upper  surface. 
These  slag  blocks  are  molded  on  tlie  flat,  in  the  same  manner  as 
ordinary  red  brick;  and  after  leveling  off  the  pile  of  dampened 
sand  to  form  a  smooth  and  horizontal  bed,  the  wooden  blocks — 
some  twelve  in  number  on  each  side  of  the  skimming  door — are 
arranged  in  a  double  row,  four  inches  apart  between  blocks,  and 
the  same  distance  between  the  two  parallel  rows. 

Besides  the  ordinary  deep  excavation  for  the  plate  slag,  a  second 
bed  should  be  left  on  each  side,  between  the  former  and  the  first 
brick  mold  right  and  left,  both  for  the  purpose  of  settling  any 
grains  of  metal  that  may  be  accidentally  drawn  over  during  the 
process  of  skimming,  and  to  act  as  a  regulating  reservoir  to  lessen 
•the  sudden  impulse  of  the  waves  of  shig  that  follow  each  motion 
of  the  rabble,  and  thus  to  prevent  the  destruction  of  the  very 
fragile  sand  molds.  The  entire  bed  is  constructed  on  an  inclina- 
tion of  about  one-half  inch  to  the  foot;  the  plate  slag  forming  the 

THE    ROASTING    OF    OKES    IN    LUMP   FORM.  143 

summit,  while  the  double  row  of  molds  slopes  away  from  it  iu  each 
direction  laterally.  After  the  wooden  blocks  have  been  placed  on 
this  sloping  bed  in  a  proper  horizontal  position,  and  exactly  equi- 
distant from  each  other,  as  determined  by  a  wooden  gauge,  the 
remaining  sand,  very  slightly  but  equably  dampened,  is  shoveled 
back  again,  and  carefully  trodden  and  tamped  evenly  into  all  the 
interspaces  and  around  the  outside  edges  of  the  blocks,  until  it 
reaches  the  level  of  their  upper  surface.  This  is  a  very  brief  oper- 
ation; for  it  is  not  essential  to  tamp  the  sand  very  firmly  so  long 
as  about  an  equal  degree  of  solidity  is  imparted  to  all  portions  of 
it.  A  cylinder  of  hard  wood  — 3  inches  in  diameter  and  4  inches 
long — which,  when  placed  lengthwise,  fits  exactly  between  each 
two  molds,  is  laid  upon  its  side,  and,  by  a  few  blows  of  the  mallet, 
driven  into  the  sand,  thus  when  removed  forming  a  little  gutter 
through  the  middle  of  the  partition  wall,  and  connecting  each 
pair  of  adjacent  cavities  in  such  a  manner  that  the  flow  of  slag 
through  either  entire  lateral  system  meets  with  no  impediment. 
The  wooden  blocks  are  then  removed  from  their  sand  bed  with 
the  greatest  care,  it  often  being  necessary  to  loosen  them  by  gentle 
tapping  and  other  means  familiar  to  the  experienced  molder.  The 
bed  requires  only  a  few  hours'  drying  to  fit  it  for  the  slag. 

By  the  time  the  charge  is  ready  for  skimming,  say  in  three 
hours  or  less  after  the  completion  of  the  bed  just  described,  it 
should  be  in  proper  condition,  and  the  furnace  helper,  armed  with 
a  small  rabble-shaped  hoe,  stands  beside  the  skimmer  ready  to 
turn  the  stream  of  slag  into  the  proper  molds,  remove  obstructions 
from  the  gutters,  br  ak  through  the  rapidly  forming  crust  if  indi- 
cations of  chilling  appear  on  the  surface  of  the  molten  bath,  and 
see  in  general  that  the  process  of  filling  the  molds  proceeds  in  a 
proper  manner.  As  soon  as  this  operation  is  concluded,  a  few 
shovelfuls  of  sand  should  be  thrown  over  the  surface  of  the  slabs  to 
prevent  sudden  and  unequal  chilling.  By  the  time  the  new  charge 
is  in  the  furnace  and  the  assistant  is  at  liberty  to  attend  to  his 
bricks,  they  will  usually  be  found  ready  for  removal,  though  still 
at  a  red  heat  on  the  surface  and  in  most  cases  quite  liquid  in  the 
interior.  It  is  essential  that  they  be  removed,  and  the  slight 
roughnesses  that  arise  from  the  broken  ends  corresponding  to  the 
gutters  through  which  they  were  filled  be  trimmed  off  with  a  small 
cutting  hammer  while  they  are  still  quite  hot,  as  it  is  just  at  this 
stage  that  they  possess  the  liighost  degree  of  toughness,  and  per- 
mit of  manipulations   that,    if    they    were  cool,    would   inevitably 


break  them  into  fragments.  These  slabs  are  best  removed  from 
the  furnace  by  being  loaded  upon  the  low  iron  barrow  commonly 
used  for  the  transportation  of  pigs  of  slag  and  matte.  The  loading 
is  effected  by  means  of  a  long  tive-eighths  inch  iron  rod,  bent  into 
a  hook  at  one  end,  and  the  blocks  are  then  wheeled  out  upon  the 
dump,  where  a  special  workman  trims  them  properly,  rejecting  all 
that  are  imperfect  or  already  cracked,  and, when  cool,  piles  them 
into  rows,  to  remain  until  needed.  The  most  useful  size  for  gen- 
eral purposes  has  been  found  to  be  about  8  by  10  by  20  inches,  and 
weighing  about  85  pounds;  but  by  simply  changing  the  form  of 
the  pattern,  they  may  be  produced  of  any  desired  shape  or  i-he, 
although  experience  has  shown  that  it  is  not  economy  to  attempt 
the  manufacture  of  very  thin  slabs,  or  of  any  weiglit  below  45 
pounds.  The  immense  value  of  this  building  material,  produced 
from  an  otherwise  worthless  substance  and  obtainable  in  rectangular 
shape  for  plain  walls  and  foundations,  in  wedge  shape  for  arches 
and  for  forming  a  circle  in  walling  wells  and  for  many  other  daily 
needs,  can  be  fully  appreciated  only  by  those  who  have  had  occa- 
sion to  build  in  a  country  where  rock  was  unobtainable  and  brick 
poor  and  expensive. 

The  slow  cooling,  or  tempering,  of  slag  will  greatly  increase  its 
toughness  and  strength,  but  it  is  only  in  late  years  that  this  method 
has  been  applied  to  the  manufacture  of  slag  brick  from  the  basic, 
ferruginous  slags  of  ordinary  blast-furnace  work.  The  following 
description,  with  illustrations,  is  taken  from  Dr.  Egleston's  paper 
on  ''The  Manufacture  of  Slag  Brick  in  Montana,;'  The  School  uf 
Mines  Quarterly^  Vol.  XII. 

The  process  which  is  used  at  the  Parrot  Works  at  Butte,  Mon- 
tana, was  invented  by  Mr.  J.  E.  Gaylord,  of  that  company,  and  is 
interesting  because  it  allows  of  (juickly  arriving  at  the  result  with 
ordinary  labor,  and  is  applicable  anywhere  and  to  almost  any  slag, 
provided  it  holds  together  on  cooling,  as  almost  all  the  slags  in  the 
West  do.  It  consists  simply  in  dumping  the  fluid  slag  from  the 
inside  of  the  ordinary  conical  iron  pot  into  a  cast-iron  mold, 
instead  of  allowing  it  to  get  cool  in  the  pot  and  then  throwing  it 
on  to  the  dump  heap.  This  requires  that  the  casting  yard  shall 
be  near  the  furnaces,  so  that  the  slag-pots  shall  not  have  to  be 
wheeled  too  far,  and  that  the  space  shall  be  large  enough  for  the 
men  to  work  conveniently,  and  also  space  for  the  storage  of  the 
hot  molded  slag  while  cooling.  The  plant  required  for  this  man- 
ufacture is  of  the  simplest  description,  and  the  product  available 

»     ^ 

I  -5 





for  almost  auy  biiildiug  required  about  the  works,  or,  indeed,  for 
any  ordinary  constrnctiou,  especially  for  underground  work. 

The  slag-bricks  at  these  works  are  made  by  contract,  and  are 
paid  for  at  the  rate  of  85  cents  to  ^1  per  hundred.  The  bricks 
are  12  inches  long  and  0  inches  wide  and  high.  This  has  been 
found  by  experience  to  be  the  most  convenient  size,  but  they 
might  be  made  of  any  other  size  when  it  was  desirable  to  do  so. 
The  Parrot  brick  weighs  about  55  pounds;  one  man  can  make  350 
in  a  day.  They  are  made  on  an  area  near  the  blast-furnaces.  At 
these  works  there  are  two  plants  for  making  them,  each  plant 
having  three  sets  of  apparatus  at  a  distance  of  about  30  feet  apart. 
These  three  sets  are  worked  by  one  man  in  the  day  and  one  in  the 
night  shift;  or  four  men  in  24  hours.  When  there  is  a  greater 
demand,  extra  sets  can  be  easily  set  up  or  shifts  of  eight  hours  can 
be  made.  The  apparatus  consists  of  a  set  of  cast-iron  plates, 
shown  in  the  accompanying  plan  and  elevation.  These  plates  are 
cast  in  the  shape  of  a  T  and  have  beveled  ends.  They  are  one  inch 
thick,  14  inches  long,  12  inches  wide  on  the  inside  and  14  on  the 
outside.  The  bevel  occupies  one  inch,  so  that  the  available  inside 
space  is  12  inches  long  and  6  inches  wide  and  high.  This  piece  is 
set  upon  a  series  of  bed -plates,  which  are  14  by  6  and  1^ 
inches  thick.  These  are  leveled  up  and  form  a  floor,  and  are  jux- 
taposed so  as  to  leave  their  joints  under  the  mold  frames.  The 
frames  are  placed  together,  so  as  to  form  five  molds,  so  that  the 
pointed,  beveled  ends  of  the  long  end  of  the  T  fit  into  the 
V-openings  made  by  placing  the  beveled  ends  of  the  short  ones 
together.  The  metliod  of  placing  them  is  shown  herewith.  No 
special  end  pieces  are  made  for  the  purpose  of  resisting  the  jires- 
sure,  but  two  of  the  castings  are  placed  at  the  end  for  that  purpose. 
On  the  outside  of  these  and  resting  upon  supports,  9  inches  high, 
plates  of  cast-iron  6  inches  wide  and  of  the  same  thickness  are  set, 
a  little  longer  than  the  length  of  the  five  molds.  They  are 
reached  by  an  incline  three  feet  long,  placed  as  shown  in  the  cut, 
so  that  the  wheels  of  the  slag-pot  will  run  on  them  and  be  just 
over  the  molds  below.  The  incline  is  so  gentle  that  there  is  no 
difficulty  in  pushing  the  slag-pot  up  it.  The  pot  full  of  slag  from 
the  furnaces  is  run  up  this  incline.  The  man  shoving  it  makes 
two  holes  in  the  crust,  which  has  cooled  on  the  top  while  coming 
from  the  furnace,  the  front  one  to  pour  the  slag  out  and  the  one 
on  the  other  side  behind  it  to  allow  of  tlie  flow.  He  then  tips  the 
pot  over  by  raising  the  handle  of  the  slag-wagon,  and  the  melted 

THE    ROASTIXG    OF    OKES   IN    LUMP   FORM.  147 

slag  on  the  inside  falls  into  the  molds  below  until  they  are  full. 
There  will  then  be  a  shell  of  slag  on  the  inside  of  the  pot.  This 
is  carried  to  the  dump-heaps  and  tipped  there,  and  the  pot  is  taken 
back  to  the  furnace  to  be  again  filled.  By  the  time  the  molds  in 
plant  No.  3  are  full,  the  brick  man,  who  has  just  prepared  these 
molds  and  has  watched  the  operation  of  casting,  is  ready  to  take 
to  pieces  the  bed  No.  2  previously  cast.  He  goes  there,  and  with 
a  hook  which  fits  into  the  holes  on  the  top  of  the  castings,  shown 
in  the  elevation,  pulls  out  the  irons  and  puts  them  to  one 
side,  leaving  the  hot  bricks  on  the  iron  pavement  to  cool  suffi- 
ciently to  be  handled.  When  this  is  done  he  goes  to  No.  1,  the 
bricks  of  which  have  been  cooling  and  are  ready  to  be  piled,  but 
are  still  hot.  He  takes  them  up  on  a  shovel  and  piles  them  close 
together,  making  headers  every  other  row.  He  then  i-econstructs 
the  molds  in  No.  1,  putting  the  irons,  which  are  still  hot,  in  place 
by  means  of  che  hook,  washes  them  with  clay  water,  and  by  this 
time  the  bricks  ot  No.  2  are  ready  to  be  piled.  He  first  goes  to 
No.  3,  pulls  out  the  irons  and  then  piles  the  bricks  of  No.  2,  and 
by  this  time  fresh  slag  comes  to  No.  1,  and  so  on.  It  does  not 
take  much  more  than  ten  minutes  between  the  casting  of  one  set 
and  the  making  of  the  piles  of  the  other.  The  bricks  are  left  in 
the  pile  until  they  are  quite  cool,  by  which  time  they  are  sufficiently 
annealed  to  be  used. 

There  is  always  a  considerable  quantity  of  small  stuff,  made  bv 
the  slopping  of  the  slag.  This  is  taken  away  by  one  man  with  a 
horse  and  cart,  and  is  used  for  making  the  roads  about  the  works 
and  for  filling  either  between  masonry  or  in  the  ground.  These 
bricks  are  constantly  used  about  the  works,  and  considerable  quan- 
tities of  them  are  sold  to  be  used  in  the  town.  They  are  very 
advantageous  for  construction,  as  they  require  less  mortar  than 
ordinary  bricks,  and  are  quite  as  strong  as  stone,  when  they  are 
not  liable  to  shock.  They  are  used  exclusively  in  the  construc- 
tion of  the  kilns  at  that  works,  where  they  would  last  a  very  loug 
time,  but  for  the  habit  of  cooling  down  the  hot  ore  with  water, 
which  makes  it  necessary  to  reconstruct  them  every  four  or  five 
years.  The  bricks  of  the  size  made  here  are  the  most  convenient. 
If  made  smaller  they  would  cost  too  much,  since  the  labor  would 
be  about  the  same  whatever  the  size  If  made  larger  they  would 
be  too  heavy  for  the  men  to  handle  conveniently.  They  can  be 
transported  short  distances  and  are  cheaper  and  more  easily  laid 
than  stone. 



The  skill  of  the  manufacture  is  entirely  in  keeping  the  irons 
above  ground,  moving  them  frequently  and  keeping  them  coated 
with  clay  water.  When,  as  in  some  cases,  the  molds  have  been 
permanently  fixed  in  place  and  the  slag  allowed  to  cool  in  them, 
the  cast-iron  pieces  have  beconie  useless  in  a  short  time.  At  the 
Parrot  Works,  where  the  work  is  done  carefully,  they  last  indefi- 
nitely, and  where  the  molds  are  taken  to  pieces  as  soon  as  the 
bricks  are  strong  enough  to  hold  themselves  up,  the  wear  is  inap- 
preciable. The  process  is  a  very  ingenious  and  simple  one  auO 
applicable  at  any  works  producing  slag.  The  cost  of  the  plant  is 
very  small,  the  labor  required  is  not  high-priced,  and  over  two- 
thirds  of  the  slag  is  a  source  of  a  sjuall  profit  to  the  works,  instead 
of  being  an  incumbrance  and  a  source  of  expense. 

SCALE  K  IN.  =  1   FOOT 



_« I. 

I  I 

\     \     L 


a.a.—Ihraught  holes  oonnecNnn 
with  Flue  in  sideiralls. 

b.b.— Flue  holes  into  Main 


I.I.  Lzr; 


I      I      I 

_flf    a. 


Fig.  20. 

To  return  to  the  roasting  stalls.  Assuming  that  they  are  to  be 
built  of  the  material  just  described,  and  without  any  iron-work 
for  anchoring,  and  that  each  stall  is  to  burn  a  charge  of  20  tons 
and  be  again  cleared  out  in  10  days,  thus  furnishing  2  tons  a  day, 
it  will  require  some  50  stalls  to  furnish  100  tons  of  ore  a  day, 
allowing  some  12  per  cent,  in  excess  of  the  needful  capncity  to 
permit  of  repairs.  The  weight  of  ore  as  brought  to  the  stalls,  and 
not  as  tahen  from  them,  is  counted:  its  loss  during  the  process  of 
calcination  depends  upon  the  quality  and  amount  of  sulphides 
present,  and  frequently  reaches  15  per  cent.,  though  a  considera- 
ble portion  of  the  loss  in  weight  due  to  <-he  elimination  of  the 
sulphur  is  offset  by  the  gain  in  oxygen. 

THE    ROASTING    OF    ORES    IN    LUMP    FORM.  140 

Such  a  battery  of  stalls  should  always  be  built  in  a  double  row, 
back  to  back,  each  lateral  wall  serving  as  the  division  between 
the  two  adjacent  partitions,  while  the  unbroken  rear  walls  form 
the  sides  of  the  main  flue,  a  space  of  at  least  two  feet  being  left 
between  them,  which  simply  requires  a  4-inch  brick  arch  to  form 
the  main  flue  for  the  entire  system.  This  also  constitutes  a  foun- 
dation on  which,  after  a  little  leveling  up  with  earth,  to  prevent 
the  sleepers  from  being  affected  by  the  heated  masonry  below,  the 
narrow  railroad  is  laid  on  which  the  ore  for  roasting  is  brought  to 
any  part  of  a  given  stall  by  means  of  the  turn-plate  and  movable 
rails,  explained  in  the  chapter  on  "  Heap-Roasting."  A  double  row 
of  28  stalls  (56  in  all)  should  have  a  flue  at  least  2  by  4  feet  for 
the  third  of  the  number  nearest  the  chimney,  which  may  be  re- 
duced to  2  by  3  feet  for  the  middle,  and  2  by  2^  feet  for  the  end 
third,  if  any  saving  can  be  effected  thereby.  The  two  long  rear 
walls,  enclosing  the  main  flue,  should  be  32  inches  thick — once 
and  a  half  the  length  of  a  slag-brick — with  proper  allowance  for 
mortar  and  irregularities,  and  should  be  laid  solely  in  clay  mortar, 
which  designation  throughout  this  entire  work  may  be  understood 
to  mean  merely  common  sticky  mud,  such  as  is  employed  for  mak- 
ing a  poor  quality  of  red  brick.  If  ordinary  clay  be  accessi))le,  it 
may  be  mixed  with  sand  in  such  proportions  as  to  slip  easily  from 
the  trowel;  otherwise,  any  ordinary  sticky  mud  may  be  used,  and 
will  be  found  to  form  perfectly  satisfactory  material  for  laying  all 
mason-work  that  is  to  be  exposed  to  sulphur  fumes  and  a  heat  not 
exceeding  a  dull  red. 

The  fact  that  lime  mortar  is  totally  unadapted  to  ordinary  met- 
allurgical uses,  although  doubtless  universally  known,  is  for  some 
unaccountable  reason  frequently  not  acted  upon,  and  the  result  in 
most  cases  is  the  rapid  and  total  destruction  of  the  furnace-arch, 
chimney,  flue,  or  whatever  structure  may  happen  to  have  been 
put  together  with  such  unfit  material.  The  acid  vapors  immedi- 
ately form  a  sulphate  with  the  lime  present  in  the  mortar,  and 
this,  absorbing  water,  becomes  gypsum  and  crystallizes,  expanding 
with  great  force,  breaking  the  Joints,  and  soon  crumbles  and 
washes  away.  It  is  quite  proper  to  use  lime  mortar  in  such  por- 
tions of  the  structure  as  are  free  from  contact  with  heat  and  sul- 
phurous gases,  and  yet  require  unusual  strength,  which  cannot  be 
obtained  with  the  clay  substitute.  Such,  for  instance,  as  in  the 
construction  of  chimneys  for  metallurgical  purposes,  where  tlie 
best  results  can  only  be  obtained  by  the  employment  of  both  of 

THE    ROASTING    OF    ORES    IN    LUMP    FORM.  151 

these  substauces:  lime  mortar  for  the  outside  Avork,  while  the 
common  chiy  mud  is  merely  used  for  the  inside  layer,  and  the 
joints  thoroughly  protected  against  any  invasion  of  the  sulphur  gases 
by  plastering  the  whole  interior  with  a  thin  coating  of  clay  mortar, 
tempered  with  sand  to  such  an  extent  that  it  will  not  crack  and  fall 
off  in  sheets.  Further  reference  will  be  made  to  this  point  in 
speaking  of  "Furnace  Building."  The  constant  and  flagrant  viola- 
tion of  this  law  is  a  sufficient  reason  for  its  frequent  reiteration. 
A  recent  example  suggests  itself,  where  the  arches  of  a  number  of 
very  expensive  and  nearly  new  calcining-furnaces  had  fallen  in, 
causing  a  very  heavy  loss.  A  conversation  with  the  mason  who 
built  them  brought  out  tiie  fact  that  they  were  constructed  with 
lime  mortar,  he  having  had  no  orders  to  the  contrary. 

The  size  of  the  stall  is  somewhat  dependent  upon  the  quality  of 
the  ore  to  be  roasted,  a  highly  siliceous  ore  with  a  comparatively 
low  percentage  of  sulphur  permitting  a  much  wider  and  higher 
stall  than  an  ore  with  little  gangue,  and  especially  than  one  contain- 
ing a  considerable  portion  of  iron  pyrites,  in  which  case  extensive 
and  unavoidable  sintering  will  follow  any  attempt  at  increasing 
the  size  of  the  stall.  A  safe  size  for  an  average  ore,  containing  a 
moderate  amount  of  pyrite  and  demanding  careful  handling,  is  8 
feet  in  length  by  6  feet  in  height  by  8  feet  in  width.  It  is  best  to 
build  the  lateral  walls  of  the  same  tliickuess  as  the  rear  division, 
the  increased  stability  and  durability  of  the  entire  structure  well 
repaying  the  slight  additional  expense  in  labor  and  material.  The 
bottom  should  be  paved  with  the  same  slabs  placed  flatwise  and 
exactly  reversed  from  the  position  in  which  they  lay  when  formed; 
their  upper  surface  now  going  downward,  while  their  original 
lower  surface,  which  should  be  perfectly  smooth  and  level,  noAv 
comes  upward.  The  connection  with  the  main  flue  is  effected  by 
means  of  8  or  10  small  rectangular  openings — 3  by  0  inches — in 
the  rear  wall,  in  two  or  more  rows,  and  at  a  considerable  distance 
from  the  ground.  These  are  kept  tightly  closed  by  means  of  a 
bunch  of  old  rags  or  a  ball  of  clay,  when  there  is  no  occasion  for 
their  remaining  open;  otherwise,  the  draught  of  the  entire  system 
might  suffer. 

The  only  air  admitted  to  these  stalls  originally,  at  the  Parrot 
works,  came  through  sucli  interstices  as  were  left  in  roughly 
building  up  the  temporary  frnnt  wall;  but  experiments  led  to  the 
addition  of  some  4  or  6  similar  openings  in  each  lateral  wall,  which 
did  not  communicate  with  the  main   culvert,  but  connected  with: 


the  outside  air  by  means  of  a  small  flue  running  longitudinally 
through  each  division  wall,  though  not  extending  so  far  as  the 
central  passage  This  innovation  has  been  followed  by  a  decided 
improvement  in  the  oxidation  of  the  ore  and  a  great  diminution  in 
the  amount  of  matte  produced.  An  essential  precaution  in  the 
management  of  these  stalls  is  to  maintain  a  thick  coat  of  clay  plas- 
tering over  their  entire  interior  surface,  by  which  the  heated  ore 
is  kept  from  sticking  to  the  walls  and  causing  the  rapid  destruc- 
tion of  the  mason-work.  A  few  moments'  attention  to  the  empty 
structure  after  each  operation  will  keep  the  plastering  intact  and 
greatly  lessen  the  cost  of  repairs.  As  the  entire  success  of  this 
process  depends  upon  the  strength  and  regularity  of  the  draught, 
a  stack  of  considerable  size  and  height  is  essential. 

A  battery  of  56  stalls,  as  described,  requires  at  sea-level  a  chim- 
ney 75  feet  high,  and  with  an  internal  area  of  at  least  I'i  square 
feet,  as  will  be  further  explained  in  the  chapter  on  the  construc- 
tion of  calciniug-furnaces.  Any  economy  in  the  direction  of 
diminishing  the  size  of  this  important  adjunct  will  be  immediately 
noticed  in  the  lengthening  of  the  roasting  process,  and  may  reduce 
the  capacity  of  the  stalls  to  an  incredible  degree.  The  draught  is 
regulated  by  means  of  a  sheet-iron  damper  hung  in  the  main  flue, 
close  to  its  junction  with  the  chimney,  while  the  same  office  is 
accomplished  for  individual  stalls  by  partially  filling  the  draught- 
holes  in  the  rear  wall  witli  bits  of  bricks  or  balls  of  clay.  In  no 
portion  of  the  process  are  the  skill  and  care  of  the  roasting  fore- 
man better  displayed  than  in  his  management  of  the  draught, 
which  must  be  varied  according  to  the  season  and  temperature  of 
the  air,  as  well  as  with  the  changing  character  of  the  ore. 

As  already  intimated,  a  stall  of  the  size  and  pattern  just  described 
will  hold  5U  to  30  tons  of  pyritous  ore,  which  should  be  kindled 
with  the  very  smallest  possible  quantity  of  wood  that  will  set  it 
thoroughly  on  fire.  This  is  essential  for  a  far  more  important 
reason  than  the  mere  saving  in  fuel;  for  the  slightest  increase  in 
tlie  contents  of  the  bed  of  wood  on  which  the  rock  is  heaped  will, 
with  pyritous  or  otherwise  easily  fusible  ores,  cause  an  amount  of 
sintering  and  a  formation  of  matte  entirely  disproportionate  to  the 
cause.  Repeated  trials  can  alone  determine  the  various  minuti* 
of  this  description  essential  to  the  best  possible  results  with  the 
material  under  treatment;  but,  in  most  cases,  where  the  ore  is  at 
all  pyritous,  good  sound  wood  will  be  found  to  produce  too  fierce 
a  heat  foi*  the  purpose,  and  recourse  must  be  had  to  decayed  wood, 

THE    ROASTING    OF    ORES    IN    LUMP   FORM.  153 

which  can  usually  be  obtained  at  from  one-half  to  two-thirds  of 
the  price  of  the  sound  fuel.  For  an  ore  containing  30  per  cent, 
sulphur  and,  say  25  per  cent,  silica,  25  cubic  feet  of  rotten  wood, 
or  about  one-fifth  of  a  cord,  will  be  found  ample;  but  this  small 
proportion  of  fuel — only  one-hundredth  of  a  cord  to  the  ton — 
must  be  utilized  in  a  proper  manner,  and  with  the  most  rigid 
economy  and  exactitude,  or  the  heap  will  miss  fire  completely, 
doubling  the  cost  of  the  operation,  as  well  as  interfering  with  the 
estimated  production  of  the  plant.  A  quarter  of  an  hour  spent  in 
watching  the  manipulations  of  an  experienced  roaster  is  better 
than  pages  of  description,  though  the  operation  of  preparing  a 
stall  for  its  ore  charge  is  far  from  complicated. 

After  seeing  that  the  layer  of  clay  on  the  enclosing  walls  is  re- 
newed with  the  plastering-trowel  where  necessary,  and  that  the 
draught-holes  are  open  to  the  extent  dictated  by  former  experi- 
ence, a  central  longitudinal,  and  two  lateral  flues  are  constructed 
on  the  floor  of  the  stall  out  of  large,  irregular  fragments  of  ore. 
These  are  merely  to  introduce  air  to  the  iiiterior  and  to  insure  the 
rapid  and  thorough  kindling  of  the  entire  structure.  They  are 
filled  and  surrounded  with  dry  kindling-wood,  and  the  greater 
part  of  the  fuel,  split  into  long,  thin  sticks  from  the  large  rotten 
logs  and  poles  that  are  usually  provided,  is  disposed  in  a  thin  layer 
over  the  bottom  of  the  stall,  the  amount  slightly  increasing  toward 
each  side.  The  structure  is  now  filled  with  coarse  ore,  and  the 
ragging  distributed  throughout  the  entire  contents  rather  than 
concentrated  in  a  considerable  layer  merely  upon  the  surface.  As 
the  stall  becomes  gradually  filled,  single  small  sticks  of  wood  are 
placed  between  the  ore  and  the  lateral  and  back  walls;  while  be- 
tween the  contents  of  the  stall  and  the  front  wall,  which  is  built 
up  with  large  lumps  of  ore  or  stall  matte,  a  considerable  quantity 
of  light  wood  is  introduced  to  insure  the  thorough  desulphuriza- 
tion  of  the  anterior  surface.  A  single  carload  of  ragging  is  spread 
on  top  of  the  coarse  ore,  and  upon  this  a  three-inch  layer  of  shav- 
ings, bark,  and  chips  is  placed  as  a  bed  for  about  one  and  a  half 
tons  of  raw  fines,  which,  if  disposed  in  the  exact  manner  indicated, 
and  covered  closely  with  a  well-roasted  ore  from  a  contiguous 
stall,  will  be  thoroughly  desulphurized,  and  the  covering  layer 
itself  being  in  a  well  calcined  condition,  the  entire  contents,  after 
burning,  may  be  passed  on  to  the  next  operation.  Mr.  R.  Pearce, 
of  Argo,  uses  with  great  advantage  a  sheet-iron  cover  over  the  top 
of  his  stalls,  luted  tightly  with  clay  to  the  walls  on  which  it  rests. 


It  is  ouly  by  employing  great  care,  and  after  repeated  trials, 
that  the  requisite  skill  will  be  attained  to  thoroughly  calcine  the 
large  proportion  of  fines  just  indicated;  but  when  one  reflects  that 
it  amounts  to  some  7  per  cent,  of  the  entire  ore,  and  jierhaps 
one-half  of  the  total  fines  produced,  it  will  be  seen  that  the  result 
is  worthy  of  any  pains  that  can  be  expended  on  it.  The  large 
pieces  of  raw  ore  that  are  employed  in  building  the  flues  and  front 
wall  become  gradually  oxidized  upon  the  surface,  and  slowly  crum- 
ble away  and  mix  with  the  finished  product  until  they  totally  dis- 
appear and  are  replaced  by  fresh  pieces.  When  the  ore  is  to  be 
removed,  the  front  wall  is  taken  down,  and  the  lumps  of  ore  from 
it  are  piled  out  of  the  way,  to  be  again  used  for  the  same  purpose. 

The  stall  should  be  fired  at  night,  as  the  smoke  is  so  dense  dur- 
ing the  first  few  hours,  and  the  draught  so  sluggish,  that  only  a 
small  part  of  the  fumes  find  their  way  into  the  proper  channel; 
but  by  the  time  the  wood  is  consumed,  the  entire  structure  has 
become  so  much  warmer  as  greatly  to  improve  the  draught.  The 
sulphur  and  other  products  of  volatilization  and  "sweating" — 
alluded  to  in  describing  the  management  of  roast-heaps — form  a  sort 
of  crust  upon  the  surface,  and  seal  all  interstices  connecting  with 
the  atmosphere,  and  force  nearly  all  fumes  to  pass  into  the  flue, 
thus  greatly  abating  a  nuisance.  For  the  first  twenty-four  hours, 
the  fire  is  confined  to  those  portions  of  the  ore  that  were  in  imme- 
diate contact  with  the  fuel.  The  process  of  oxidation  advances 
very  "rapidly,  and  by  the  close  of  the  second  day  it  is  hardly  possi- 
ble to  bear  the  hand  upon  the  middle  of  the  upper  surface  of  the 
stall,  showing  that  at  least  one  half  the  contents  is  already  in 
combustion.  By  the  end  of  the  fourth  day  a  similar  degree  of 
temperature  may  be  felt  upon  the  upper  surface,  at  the  very  back 
of  the  stall,  proving  that  the  process  has  by  that  time  invaded  the 
entire  length  and  breadth  of  the  stall,  though  considerable  time  is 
still  necessary  for  its  thorough  completion. 

The  successful  progress  of  the  process  is  clearly  marked  by  the 
great  rise  in  height  of  the  entire  contents,  gaining  some  three 
inches  in  a  single  day,  and  frequently  ascending  some  12  inches 
above  the  level  of  the  walls  at  which  it  stood  at  the  beginning  of 
the  operation,  aside  from  the  free  space  left  to  be  filled  out  with 
ore  from  the  disappearance  of  the  fuel,  amounting  to  some  25 
cubic  feet.  This  striking  phenomenon,  unfamiliar  to  those  accus- 
tomed oulv  to  heap-roasting,  where  a  settling  rather  than  a  rising 
of  the  entire  mass  occurs,  is  simply  due  to  the  fact  that,  in  all 

THE    ROASTING    OF    ORES   IN    LUMP   FORM.  155 

cases  of  oxidizing  roasting,  a  greater  or  less,  though  always  very 
marked,  increase  in  bulk  occurs  from  the  swelling  and  Assuring  of 
the  oxidized  ore.  The  contents  of  the  roast-heap,  being  perfectly 
free  and  unconfined,  simply  spread  out  laterally,  while  the  con- 
sumption of  the  thick  bed  of  fuel  on  which  it  rests  detracts  consid- 
erably from  its  height.  The  walls  of  the  stall,  however,  enclose 
the  ore  in  a  rigid  grasp,  making  it  absolutely  necessary  that  any 
increase  in  bulk,  beyond  that  very  slight  amount  necessary  to  re- 
place the  space  occupied  by  the  fuel,  should  take  place  vertically. 
In  a  badly  burned  stall,  where  extensive  sintering  has  taken  place, 
and  a  sufficient  amount  of  the  sulphides  has  melted  into  a  solid 
mass  to  cause  a  decided  diminution  in  bulk  instead  of  an  increase, 
the  occurrence  of  crater-like  depressions  in  the  surface  of  the  ore 
is  positive  evidence  of  such  local  fusions.  That  the  pressure  re- 
sulting from  the  increase  in  bulk  is  something  quite  tangible,  may 
be  inferred  from  the  frequent  pushing  outward,  or  even  overturn- 
ing of  the  heavy  lateral  walls  of  a  stall,  provided  one  or  the  other 
of  its  contiguous  compartments  is  either  empty  or  unbraced,  while 
the  temporary  front  wall  would  inevitably  be  thrown  down  within 
the  first  day  after  kindling  if  not  strongly  supported  by  timbers. 

The  length  of  time  necessary  for  the  process  under  consideration 
is  another  uncertain  factor.  If  the  stall  be  left  undisturberl,  it 
will  usually  burn  quietly  for  a  period  of  twelve  days,  demanding 
little  or  no  attention  beyond  an  occasional  shovelful  of  covering  if 
heating  too  fiercely  at  any  one  point,  and  requiring  about  three 
<lays  additional  to  cool  sufficiently  to  remove  with  comfort;  but, 
under  ordinary  everyday  circumstances,  no  such  moderation  can 
be  practised,  and  the  period  of  each  operation  can  be  curtailed, 
without  any  especial  damage,  to  one-half  this  time.  To  accom- 
plish this  without  detriment  to  the  i^rocess  of  desulphurizatiou, 
the  following  precautions  must  be  ado23ted:  As  soon  as  the  ante- 
rior surface  of  the  ore  is  so  cool  as  to  impart  no  disagreeable  sensa- 
tion to  the  hand,  the  temporary  front  wall  should  be  removed,  the 
natural  adhesion  common  to  all  sulphureted  ores  when  roasted  in 
lumps  preventing  the  caving  of  the  vertical  ore  face,  which  should 
))e  most  carefully  attacked  with  pick  and  shovel,  every  precaution 
being  taken  not  to  penetrate  beyond  the  line  of  comparative  cool- 
ing, and  only  so  much  ore  being  removed  at  any  one  operation  as 
is  consistent  with  the  uninterrupted  progress  of  the  roasting  in 
the  mass  behind.  At  least  six  or  eight  inches  of  ore  should  be  left 
between   the  outer  air  and   the  line  of  fire,  and  any  sudden  eleva- 


tion  of  the  surface  temperature,  as  well  as  increaseJ  diflQculiy  in 
detaching  the  ore  from  the  face  on  which  work  is  prosecuted,  is  a 
sign  to  stop.  To  illustrate  the  ease  with  which  the  contents  of  a 
well-burned  stall  can  be  handled,  the  entire  charge  of  ore  from 
such  a  stall  can  be  removed  with  nothing  stronger  than  a  shingle. 

The  first  carload  is  usually  taken  from  the  stall  at  the  close  of 
the  fourth  day,  and  the  amount  capable  of  removal  may  be  rapidly 
increased,  until  in  seven  days  more  the  compartment  is  again 

By  this  careful  method  of  constant  and  systematic  slicing,  some 
two  or  three  tons  of  well-burned  ore  may  be  taken  daily  from  each 
of  -40  or  50  stalls,  and  the  capacity  of  the  roasting  plant  rendered 
more  than  double  what  it  would  be  if  they  were  left  untouched  for 
the  time  necessary  for  their  complete  desulphurization  and  cooling; 
while  the  process  of  oxidation  does  not  suUer  in  the  slightest  de- 
gree if  the  precautions  Just  enumerated  are  adhered  to. 

In  the  case  of  ores  coutaiuing  arsenical  pyrites,  or,  indeed,  in 
the  presence  of  any  form  of  arsenical  or  antimouial  combinations, 
a  considerable  proportion  of  the  same  that  would  otherwise  go  into 
the  next  operation  in  the  shape  of  antimonates  and  arsenates  may 
be  volatilized  and  completely  dispersed  by  the  admixture  of  chips, 
small  coal,  brushwood,or  other  carconaceous  materials,  which,  as 
in  heap-roasting,  exercise  a  powerful  reducing  influence  upon  the 
products  of  oxidation  just  mentioned,  and  volatilize  them  in  a 
metallic  form.  This  simple  precaution  is  of  much  greater  value 
in  the  calcination  of  similar  compounds  in  a  pulverized  condition 
in  furnaces,  where  the  different  periods  of  oxidation  and  reduction 
are  under  the  control  of  the  operator,  and  can  be  made  to  follow 
each  other  in  the  manner  most  conducive  to  the  object  in  view; 
but  even  in  the  rude  process  under  consideration,  experience  has 
shown,  in  many  cases,  that  a  decided  improvement  in  the  grade  of 
copper  has  resulted  from  this  device,  the  simplicity  and  economy 
of  which  are  among  its  strongest  recommendations. 

The  results  obtained  in  stall-roasting  vary  little  as  compared 
with  those  from  burning  in  heaps.  On  the  whole,  it  is  not  quite 
so  easy  to  prevent  the  formation  of  matte  in  the  former  practice, 
nor  do  average  and  oft-repeated  examinations  show  quite  as  good 
results  in  the  elimination  of  the  sulphur. 

As  circumstances  may  arise  where  it  becomes  the  duty  of  the 
constructing  metallurgist  to  decide  between  these  two  systems,  to 
the  positive  exclusion  of  all  methods  involving  the  pulverization  of 

THE    ROASTING    OF    OEES   IN"   LUMP    FORM.  157 

the  ore,  aud  to  give  his  reasons  for  and  against  each  method,  that 
his  employers  may  also  have  some  idea  of  the  matter  on  which  to 
base  their  advice  or  to  rest  the  confirmation  of  his  decision,  it  will 
be  well  to  concisely  review  the  comparative  advantages  and  draw- 
backs of  heap  and  stall-roasting.* 

The  first  and  most  obvious  advantage  of  the  system  of  heap- 
roasting  is  the  apparent  cheapness  and  simplicity  of  the  plant, 
only  a  level  area  being  required,  without  furnaces,  flues,  stacks,  or 
other  expensive  appurtenances. 

The  extreme  simplicity  of  the  method  and  the  very  satisfactory 
results  obtained  under  proper  management  also  speak  in  its  favor; 
but  further  than  this  no  arguments  can  be  advanced  in  support 
of  the  process. 

Even  the  economy  in  first  cost  of  plant  will  be  found  more  ap- 
parent than  real,  when  the  expense  of  the  trestle-work  and  track, 
as  well  as  the  establishment  of  the  different  grades  between  spalling- 
shed,  roast-yard,  and  smelting-liouse  levels  are  considered,  and  no 
one  will  deny  the  absolute  necessity  for  such  an  arrangement  if 
work  on  a  large  scale  is  to  be  prosecuted  with  any  degree  of 

A  careful  comparative  calculation  of  costs,  corrected  by  the 
results  of  actual  work,  shows  that,  under  ordinary  circumstances, 
the  difference  in  cost  between  the  two  plans  under  consideration  is 
too  trifling  to  have  much  weight  in  the  choice  of  methods,  and 
may  even  be  on  the  side  of  the  stalls  in  cases  where  the  natural 
conformation  of  the  land  is  unfavorable  for  the  establishment  of 
the  terraces  necessary  for  cheap  heap-roasting. 

A  far  more  important  reason  for  the  adoption  of  the  stall  system 
is  the  great  saving  in  time,  by  which  the  delay  incidental  to  the 
cruder  process  of  calcination  is  diminished  by  at  least  80  per  cent. 

In  works  of  large  capacity,  this  becomes  a  question  of  vital  im- 
portance; for  few  smelting  companies  are  so  amply  provided  with 
capital  as  to  carry  a  constant  stock  of  some  10,000  to  50,000  tons  of 
ore,  representing  a  money  value  of  several  hundred  thousand  dol- 
lars, which  is  not  at  all  an  extravagant  estimate  for  works  of  the 
capacity  under  consideration.     The  circumstance  that  this  amount 

*  See  article  on  "  The  Mines  and  Smelting-Works  of  Butte  City,"  bv  the 
author,  in  the  United  States  publication  on  Mineral  Resources  (by  A.  Will- 
iams, Jr.,  1885).  The  third  method  of  roasting  lump  ore — that  is,  in  contin- 
uous kilns — is  only  suited  to  certain  peculiar  conditions,  and  need  not  be 
considered  when  comparing  the  other  two  systems. 


may  be  reduced  to  a  sum  uot  exceeding  one-fifth  of  the  above  by 
the  substitution  of  the  quicker  method  of  calcination  is  a  weighty 
argument  for  its  adoption. 

A  still  further  advantage  may  be  claimed  for  them  in  the  con- 
centration of  all  noxious  fumes  into  a  single  flue,  and  their  dis- 
charge into  the  atmosphere  at  such  an  elevation  as  to  insure  tiieir 
gradual  diffusion  and  dispersion  without  annoyance  or  damage. 
This  is  a  great  boon  to  the  surrounding  country,  and  more  espe- 
cially to  the  workmen  employed  in  the  process  of  roasting,  as  any 
one  familiar  with  the  atmosphere  of  an  establishment  where  heap- 
roasting  is  practised  can  testify. 

Still  further  may  be  mentioned  the  considerable  saving  effected 
by  the  thorough  roasting  of  the  entire  contents  of  the  stall,  in- 
cluding even  the  fine  covering  material,  all  of  which  is  in  condi- 
tion for  the  sncceeding  operation;  whereas,  in  the  case  of  heap- 
roasting,  at  least  10  per  cent,  of  the  entire  stock  requires  a  second 
handling.  Here  may  also  be  considered  the  serious  losses  of  metal 
from  wind,  rain,  and  other  atmospheric  causes,  which,  although 
not  entirely  obviated  by  the  eni])loyment  of  stalls,  are  at  least 
greatly  lessened;  the  saving  in  a  certain  plant  of  moderate  capacity 
amounting  in  a  single  year,  according  to  the  author's  calculations, 
to  more  than  sufficient  to  cover  the  entire  cost  of  erecting  the 

But  the  most  important  advantage  possessed  by  stall-roasting 
over  heap-roasting  in  an  ordinarily  moist  climate — if  the  process 
be  carried  on  in  the  open  air — is  the  prevention  of  loss  by  leaching. 

The  cost  of  stall  roasting  will  not  vary  far  from  50  cents  per 
ton  of  ore,  with  the  same  prices  as  are  assumed  in  the  estimate  for 

The  cost  of  a  battery  of  56  stalls,  built  in  the  manner  recom- 
mended and  reduced  to  the  standard  table  of  price  adhered  to 
throughout  this  work,  is  appended.  Their  life,  nnder  ordinary 
treatment,  will  not  exceed  six  years,  at  the  expiration  of  whicli 
time  they  will  be  found  in  such  a  condition  as  to  demand  complete 
rebuilding,  although,  of  course,  the  stack  will  outlast  several 
generations  of  stalls. 


This  being  the  first  estimate  yet  given  pertaining  to  the  con- 
struction of  any  considerable  portion  of  a  smelting  plant,  the 
quickest  and  most  convenient  method  of  arriving  at  the  desired 

THE   BOASTING   OF   ORES   IN"    LUMP   FORM.  159 

result  will  be  presented  a  little  more  in  detail  than  may  be  consid 
ered  necessary  in  subsequent  calculations. 

The  total  expense  of  the  finished  stalls  may  be  conveniently 
divided  into  the  following  heads: 

1.  Excavation  for  foundations, 

2.  Cost  of  slag-brick,  clay,  and  other  building  materials,  deliv- 
ered on  the  ground. 

3.  Labor  in  building  the  stalls. 

4.  Total  expense  of  the  railroads  belonging  to  this  part  of  the 

5.  Miscellaneous  expenses  and  superintendence. 

The  actual  expense  of  building  a  plant  of  this  description  will 
almost  invariably  be  found  much  greater  than  the  most  carefully 
prepared  estimates  would  indicate,  unless  the  figures  were  made 
by  a  man  of  long  experience  in  these  matters.  The  value  of  the 
numerous  estimates  of  cost  and  expense  contained  in  these  pages 
is  principally  due  to  the  fact  that  they  are,  almost  without  excep- 
tion, taken  from  the  results  of  actual  work,  executed  under  the 
superintendence  of  the  author.  They  may,  consequently,  lay 
claim  to  a  usefulness  and  reliability  that  the  most  carefully  pre- 
pared estimates  of  cost  would  not  possess  unless  derived  from,  or  at 
least  corrected  by,  a  long  and  thorough  personal  experience  in  such 

To  prepare  the  foundations  for  the  required  number  of  stalls, 
assuming  the  ground  to  be  comparatively  level,  will  require  about 
60  days'  labor,  aside  from  the  removal  of  the  earth.  This  allows 
for  an  8-inch  pavement,  and  for  an  extension  of  the  foundation 
walls  about  two  feet  under  ground. 

1 .     Excavation  for  foundations: 

Labor,  60  days  at  $1.50 $90.00 

Removing  tbe  excavated  material 35.00 

Superintendence  and  miscellaneous  extras 32.00 

Total $157.00 

In  order  to  estimate  the  amount  of  building  material  required, 
it  is  essential  to  determine  the  cubic  contents  of  all  the  walls  in- 
closing the  56  stalls,  28  in  each  row.  The  stalls  being  6^  feet 
wide,  and  all  walls  being  32  inches  thick,  it  will  be  seen  that  the 
entire  length  of  the  two  main  rear  walls  is  520  feet,  to  which  must 
be  added  the  aggregated  length  of  the  58  partition  walls,  each  8 
feet  long,  or  464  feet,  making  a  grand  total  length  of  984  feet. 


This  wall  beiog  6  feet  high  and  32  inches  thick  contains  in  round 
numbers  15,700  cubic  feet.  To  this  must  still  be  added  about 
one-tliird,  to  allow  for  the  foundation  walls,  and  also  the  necessary 
amount  of  slabs  for  paving  tlie  stalls.     The  details  are  as  follows: 

Main  walls 15,700  slag-brick. 

Foundation    walls 5,250         " 

Paving 2,080 

Total 23,030 

As  these  slabs  are  8  b\'  10  by  20  inches,  they  contain  very  nearly 
a  cubic  foot  each,  and  when  the  very  coarse  joints  that  they  form 
are  also  considered,  it  will  be  found  that  their  customary  rating  of 
a  cubic  foot  each  will  be  perfectly  safe.  They  are  laid  entirely  in 
ordinary  clayey  loam,  which  may  he  found  almost  everywhere,  and 
which,  if  too  sticky  to  leave  the  trowel,  will  be  greatly  improved 
liy  the  addition  of  one-fourtb  or  more  of  sand,  or  even  sandy  loam. 
At  our  standard  of  prices,  81  per  ton  will  be  ample  for  such  mate- 
rial, and  will  lay  one  hundred  brick.  The  cost  of  the  slag-brick 
has  been  placed  at  two  cents  on  the  ground,  as  their  delivery  is  at 
least  as  expensive  as  their  manufacture.  The  sum  mentioned, 
that  is,  two  cents  apiece  delivered,  or  one  cent  at  the  furnace,  will 
cover  the  cost  of  making  and  trimming,  and  leave  enough  margin 
to  occasionally  replace  the  pattern  blocks  and  other  material  neces- 
sary for  their  production. 

2.   Cost  of  mater iah  for  mason-work. 

23,000  slag-brick,  at  2  cents $460.00 

235  tons  clay,  at  $1 2.35.00 

Mortar-boxes,  bods,  screens,  etc 45.00 

Total $740.00 

The  persons  employed  for  this  work  should  on  no  account  be  the 
regular,  high-priced  brickmasons,  as  these  fare  but  badly  in  han- 
dling the  heavy,  brittle  slabs,  and  neither  like  the  work  nor  are 
able  to  earn  the  large  wages  that  they  invariably  demand  and  re- 
ceive. The  proper  mechanics  for  this  work  are  what  are  popularly 
known  as  "country  stonemasons,"  whose  apprenticeship  at  build- 
ing stone  walls,  underpinning  barns  and  houses,  etc.,  has  exactly 
prepared  them  for  handling  such  rough  and  heavy  material  as  that 
under  discussion. 

Experience  in  this  particular  kind  of  construction  has  shown 
that  the  most  advantageous  distribution  of  the  force  is  to  provide 

THE    ROASTIXG    OF    ORES    IX    LUMl'    FORM.  161 

each  stonemason  with  two  immediate  helpers,  who  assist  him  con- 
stantly, bringing  the  slab,  placing  it  in  position,  and,  in  fact, 
doing  everything  excepting  the  spreading  of  the  mortar  and  that 
last  wedging  and  chinking  that  are  of  such  vital  importance  in  the 
proper  execntion  of  work  of  this  description. 

There  are  no  hodcarriers,  as  the  slabs  are  delivered  b}'  wagons 
at  the  point  most  convenient  to  the  workmen,  and  the  mortar, 
easily  and  rapidly  manufactured  from  the  materials  already  men- 
tioned, is  brought  in  large  pails,  being  used  in  immense  quantities 
in  work  of  this  description,  although  every  crevice  should  be  well 
filled  with  small  fragments  of  rock  or  slag,  called  "spalls." 

It  has  been  found  that  each  group  of  three  men,  as  described 
above,  will  lay  on  an  average  100  slag-brick  daily,  and  not  more. 

3.  Labor  in  huUding  stalls. 

Estimate  for  layiinj  100  hvick: 

One  stonemason $3.00 

Two  laborers  at  |1.50 3.00 

Mixing  mortar  for  same 50 

Carrying  mortar  and  other  miscellaneous  labor . .       .15 

Superintendence 85 

Total  for  100 , .  , , $7.00 

Total  for  28,000  brick $1,610.00 

4.  Cost  of  Railroad  Tracks. — As  all  railroads  about  the  works 
should  be  of  the  same  gauge  and  pattern,  a  single  detailed  estimate 
will  determine  the  cost  per  foot  once  for  all.  For  tracks  of  the 
required  description,  having  a  22-inch  gauge,  and  calculated  to 
carry  a  net  load  of  1,800  pounds,  the  car  weighing  an  additional 
800  pounds,  a  good  quality  T-rail  of  not  less  than  12  pounds  to 
the  yard  should  be  selected  and  well  fastened  in  place  by  a  spike 
in  every  sleeper,  while  the  abutting  ends  of  the  rails  should  be 
firmly  secured  by  fish-plates,  tapped  for  four  t-inch  bolts,  two  to 
each  rail.  Unless  the  bolt-holes  in  both  fish-plates  and  rails  can 
be  bored  where  ordered  in  such  a  manner  that  there  shall  bo  no 
doubt  of  their  perfect  correspondence,  it  is  better  to  leave  the 
plates  blank,  and  bore  them  on  the  spot.  This  may  seem  a  slight 
matter,  but  its  neglect  sometimes  causes  serious  annoyance  and 
delay  in  outlying  districts,  and  the  boring  of  the  thin  fish  plates  is 
a  slight  task,  as  every  smelter  should  be  provided  with  a  boring- 
machine  run  by  power,  which  is  indispensable  for  sampling  pig- 
copper;  and  will  be  found  generally  useful. 


The  sleepers  are  sawed  from  the  ordinary  timber  of  the  country, 
and  may  be  conveniently  ordered  of  the  following  dimensions:  30 
inches  long,  G  inches  wide,  and  -i  inches  thick —containing  each  G 
feet,  board  measure.  They  should  be  placed  39  inches  apart  from 
center  to  center,  and  last  almost  indefinitely,  as  the  sulphate  salts 
with  which  they  become  impregnated  prevent  their  decay. 

B^or  convenience  of  calculation,  the  estimate  will  be  based  on  a 
length  of  100  yards  of  track: 

WeigJtt  of  iron  : 
200  yards  rails  at  12  pounds  =  2,400  pounds. 
Spikes,  bolts,  and  fishplates  =     115 

Total 2,515  pounds  X  3i  cents  =  $88.02 

Sleepers  : 
125  containing  6  feet  each   =   750  feet  at  $25  per  M  =   18.75 

Labor  : 

Grading,  laving  track,  ballasting,  etc $13.50 

Superintendence 6.00 

Total 19.50 

Average  allowance  for  curves  and  switches 13.6.3 

10  per  cent,  for  incidentals 14.00 


We  may  therefore  accept  the  figure  of  $1.53  per  yard,  or  51 
cents  per  foot,  as  the  cost  of  a  tram-road  of  this  description,  and 
there  being  three  lines  of  track  required  for  the  stalls,  aggregating 
a  length  of  T80  feet,  to  which  must  be  ad^^ed  100  feet  for  connec- 
tions, switches,  and  single  main  line  to  smelter,  we  have  a  total  of 
880  feet  at  51  cents,  or  *44S  80. 

Rails  for  long  curves  may  be  bent  cold;  for  short  curves,  they 
must  be  slightly  heated;  while  frogs,  points,  etc.,  require  welding, 
and  can  be  readily  constructed  in  any  ordinary  blacksmith's  forge. 

Great  care  should  be  taken  in  laying  the  track,  nor  should  the 
foreman  rest  satisfied  until  every  point,  frog,  and  guard  rail  is  in 
proper  position  and  has  the  precise  curve  necessary  for  easy  passage 
of  the  car  without  undue  friction  or  danger  of  derailment.  It  is 
scarcely  necessary  to  say  that  this  work  can  only  be  properly  and 
economically  executed  under  the  direction  of  an  experienced  rail- 
road constructor. 

5.  Miscellaneous  Expenses  and  Superintendence. — Aside  from 
the  allowance  made  in  each  department  of  the  work  for  the  above 

THE    ROASTIXG    OF    ORES    IN    LUMP   FORM.  163 

purposes,  it  will  be  found  in  practice  that  a  considerable  additional 
sum  is  required  to  cover  errors  in  construction,  blacksmith  work, 
and  various  incidentals,  as  well  as  general  superintendence, 
amounting  in  a  case  similar  to  the  above  to 

Cost  of  4-mcli  brick  arch  to  cover  main  flue 137.00 



Excavation  for  foundations $157.00 

Materials  for  mason-work 740.00 

Labor  in  building  stalls 1,610.00 

Railroads  448.80 

Miscellaneous  and  superintendence 348.00 

Grand  total $3,303.80 

Uneven  ground,  bad  weather,  and  other  unfavorable  causes  may 
increase  this  sum  to  a  considerable  extent,  but  the  figures  given 
will  be  found  safe  under  ordinary  circumstances  and  with  strictly 
judicious  and  economical  management. 

The  calcination  of  matte  in  ore  stalls  of  the  pattern  just  described 
is  by  no  means  impossible,  the  principal  difference  between  its 
treatment  and  that  of  ore  being  the  increased  quantity  of  fuel 
required — about  three  times  as  much.  A  considerable  proportion 
of  the  matte  will  be  fused  during  the  operation,  and  another  large 
fraction  scarcely  affected  by  the  process;  so  that  from  three  to 
four  burnings  are  required  to  effect  any  reasonably  perfect  desul- 

This  practice  cannot  be  recommended,  as  much  better  results 
are  obtained  by  providing  the  stalls  with  grate-bars,  and  prevent- 
ing the  radiation  of  heat  from  the  surface  by  means  of  an  arched 
brick  roof. 


This  is  a  method  well  known  in  the  Eastern  States,  and  prac- 
tised first  in  this  country,  so  far  as  any  record  can  be  found,  at 
the  old  Eevere  Copper  Works  in  Boston,  and  in  more  modern 
times  at  Copperas  Hill  in  Vermont,  and  at  the  noted  Vershire 
mine  in  the  same  State,  where  some  sixty  or  seventy  stalls  are  still 
in  use.  The  partial  suppression  of  the  excessively  disagreeable 
fumes  generated  in  the  heap-roasting  of  this  substance;  a  gain  of 
at  least  one-third  in  the  time  of  treatment — no  unimportant  item 


ill  the  handliug  of  such  valuable  material;  and  a  very  great  dimi- 
iiution  in  the  losses  caused  by  the  elements  are  the  principal  rea- 
sons for  the  selection  of  stalls  in  preference  to  heaps.  On  the 
other  hand  must  be  placed  a  heavy  investment  in  buildings  and  in 
the  stalls  themselves,  with  their  fines,  stacks,  etc.  The  mere 
grate-bars  for  a  single  matte  stall  cost  in  the  neighborhood  of  tTo, 
and  the  constant  repairs  that  are  peculiarly  necessary  in  the  case 
of  mason-work  saturated  with  the  products  of  volatilization,  and 
racked  by  the  frequent  and  extensive  fluctuations  in  temperature, 
due  to  the  ever-recurring  heating  and  cooling  of  the  interior,  ren- 
der them  a  somewhat  expensive  portion  of  the  plant,  as  will  be 
seen  in  detail  in  its  proper  place. 


The  grate-bars  being  thoroughly  cleansed  and  freed  from  all 
clinkers  and  debris  of  the  preceding  operation,  and  replaced  in 
position,  and  the  brick  walls  forming  the  sides  and  back  of  the 
stall  receiving  a  fresh  coat  of  plaster  (clay)  where  necessary,  a 
layer  of  fuel  is  placed  upon  the  grate-bars,  and  the  broken  matte 
thrown  upon  this  by  means  of  a  closely-tined  fork,  to  separate  the 
fine  stuff,  which  is  scattered  over  the  top  after  the  stall  is  filled 
with  an  average  charge  of  from  five  to  six  tons. 

The  fuel  employed  is  wood  in  4  or  6-foot  lengths,  and  split  to 
a  comparatively  uniform  size.  From  10  to  20  cubic  feet  are  used 
for  each  charge,  metal  of  low  grade,  rich  in  sulphur,  requiring  less 
fuel  than  the  higher  varieties  of  matte.  Experience  has  taught 
the  great  advantage  obtained  by  the  use  of  hard  wood,  and  too 
much  care  cannot  be  bestowed  upon  the  selection  of  the  fuel,  which 
should  be  of  the  best  quality  and  thoroughly  seasoned,  as  the  re- 
sult of  the  operation  depends  to  a  remarkable  extent  upon  the 
quality  of  the  fuel  used. 

Matte  of  any  grade,  from  the  lowest  coarse  metal  to  the  highest 
quality  of  regule,  may  be  treated  in  these  stalls  with  almost  equal 
results  as  regards  desulpliurization. 

The  stalls  are  always  covered  by  rude  sheds,  to  protect  the  brick- 
work from  the  weather,  and  should  be  paved  with  slag  blocks,  flat 
stone,  or,  much  better,  heavy  iron  plates,  as  the  constant  ham- 
mering that  it  must  undergo  during  the  spalling  of  the  matte  and 
the  breaking  of  the  huge  clinkers  that  form  an  almost  necessary 
accompaniment  of  this  process,  quickly  destroys  any  other  descrip- 
tion of  pavement.     The  results  of  desulpliurization  by  this  method 

THE    ROASTING    OF    ORES    IN    LUMP    FORM.  165 

being  uo  more  thorough  than  by  heap-roastiug,  the  same  number 
of  burnings  is  necessary  as  in  the  latter  case,  and,  owing  to  the 
difficulty  of  removing  the  heavy  clinkers  from  the  walls  and  grate- 
bars  of  these  little  furnaces,  as  well  as  the  constant  bill  of  expense 
for  repairs,  the  cost  of  the  process  is  about  the  same  as  in  heap- 
roasting.  The  almost  complete  identity  of  the  two  methods  in 
this  respect  renders  any  further  details  of  expense  unnecessary. 
The  imperfections  of  all  the  methods  of  roasting  matte  in  lump 
form,  as  well  as  the  great  waste  of  time  and  metal,  and  the  annoy- 
ance caused  by  the  fumes,  are  serious  objections,  and  it  is  only 
under  exceptional  circumstances  that  these  crude  and  dilatory 
methods  can  be  recommended.  In  nearly  all  advanced  works, 
they  have  given  place  to  the  much  more  rapid  and  perfect  method 
of  calcination  in  reverberatory  furnaces. 

The  ordinary  dimensions  of  the  stalls  in  use,  now  or  formerly, 
at  some  of  the  principal  works  in  this  country  are  as  follows: 

Width 5  feet. 

Depth  (front  to  baclv) 6  feet. 

Depth  of  ash-pit 1  foot  6  inches. 

Height  from  grate  to  spring  of  arch 4  feet  8  inches. 

Thickness  of  division  walls 1  foot  4  inches. 

Thickness  of  rear  walls  1  foot  8  inches. 

Area  of  flue  opening  in  rear  wall 160  square  inches. 

A  stall  of  this  size  will  contain  from  live  to  six  tons  of  matte, 
and  will  burn  for  four  days  at  the  first  firing,  and  for  about  three 
days  at  each  subsequent  operation. 

Where  three  burnings  take  place,  the  capacity  of  each  matte 
stall  may  be  placed  at  one-half  ton  daily,  and  the  amount  of  wood 
required  for  the  three  burnings  will  be  one-twelfth  of  a  cord  per 
ton  of  ore. 

From  the  measurements  already  given,  aided  by  the  estimates  for 
brick-work  found  in  a  succeeding  chapter,  the  cost  of  a  block  of 
sucii  covered  stalls  may  be  easily  arrived  at;  the  covering  arch 
consisting  of  a  9-inch  semicircle  of  red  bricks,  and  the  main  flue 
section  being  at  least  equal  to  the  combined  area  of  the  flues  that 
enter  it. 

The  anchoring  of  a  block  of  such  stalls  is  very  simple,  consisting 
of  longitudinal  f -inch  rods,  while  the  uprights  may  be  iron  rails 
or  stout  wooden  timbers.  Each  side  wall  should  also  be  braced  from 
front  to  back  in  the  usual  manner,  while  the  front  wall  of  the 
stall  is  a  temporary  structure  of  brick  laid  loosely  upon  the  grate- 


bars  and  braced  with  a  few  lengths  of  flat  iron.  Fire-brick  are 
ordinarily  used  for  this  purpose,  the  common  red  brick  of  which 
the  entire  permaneut  portion  of  the  structure  is  built  being  too 
light  and  fragile  to  stand  the  repeated  handlings  and  the  fluctua- 
tions of  temperature. 

Since  the  ordinary  charge  only  tills  the  stall  about  two-thirds 
full  at  the  front,  and  slopes  up  against  the  rear  wall  to  nearly  the 
height  of  the  flue  opening  near  the  top  of  the  walls,  or  even  in 
the  arched  roof,  a  large  space  exists  between  the  upper  edge  of 
the  temporary  front  retaining  wall  and  the  high  semicircular  brick 
roof.  Through  this,  the  sulphurous  fumes  and  the  products  of 
the  combustion  of  the  fuel  during  an  early  stage  of  the  process 
escape  in  such  clouds  as  to  render  the  atmosphere  of  the  slied  untit 
for  respiration.  To  partially  obviate  this  difficulty,  a  sheet-iron 
curtain,  suspended  by  wires  running  over  a  pulley  in  the  roof,  and 
furnished  with  a  counter-weight,  is  used,  and  if  properly  fitted 
and  luted  to  the  side  walls  witli  a  paste  of  stiff  clay,  is  of  great 

It  may  be  assumed  with  safety  that,  by  the  process  of  matte- 
roasting  -in  lump  form — whether  executed  in  heaps  or  covered 
stalls — from  two-thirds  to  three-fourths  of  its  original  sulphur 
contents  is  eliminated,  by  not  less  than  three  consecutive  burnings. 

THE    ROASTIXQ    OF    ORES    IX    LUMP    FORM    IN     KILNS. 

By  the  term  kiln,  as  used  here,  we  understand  a  comparatively 
small,  shaft-like  furnace,  provided  with  a  grate  or  opening  for  the 
admission  of  air  from  the  bottom,  and  connected  with  a  draught 
flue.  The  action  is  a  continuous  one,  and  the  necessary  heat  is 
derived  entirely  from  the  oxidation  of  the  sulphur  and  the  other 
constituents  of  the  ore. 

Xn  other  class  of  furnaces  has  received  greater  attention  or  been 
brought  to  a  greater  state  of  perfection;  but  it  is  as  an  adjunct  to 
the  manufacture  of  sulphuric  acid  rather  than  to  the  calcination 
of  ore  that  this  apparatus  must  be  esteemed,  and  consequently  to 
the  works  treating  on  that  subject  that  we  must  look  for  detailed 
descriptions  and  estimates  of  the  same.  The  student  is  referred 
to  Lunge's  exhaustive  work  on  "Sulphuric  Acid"  for  full  details 
of  construction  and  management. 

While  the  various  processes  of  roasting  hitherto  described  are 
suited   to   almost  every  variety  of   sulphureted  copper  ore,  and 


THE    ROASTING    OF    ORES    IN    LUMP   FORM.  167 

yield  equally  good  results  whether  the  percentage  of  sulphur  and 
copper  is  small  or  large,  a  much  closer  selection  of  material  is  in- 
dispensable for  successful  roasting  in  kilns,  and  their  range  of 
usefulness  is  restricted  to  comparatively  narrow  limits. 

This  very  question  of  selection,  however,  varies  greatly  with  the 
purpose  in  view,  and  depends  upon  whether  it  is  desired  merely  to 
desulphurize  a  given  ore  without  any  attempt  to  utilize  the  volatile 
products  of  oxidation,  or  whether  the  manufacture  of  sulphuric 
acid  is  to  be  combined  with  the  process  of  roasting. 

The  conditions  necessarily  present  before  any  pyrites  can  be 
utilized  for  the  manufacture  of  sulphuric  acid  are  of  two  kinds, 
commercial  and  technical. 

The  commercial  conditions  are  sufficiently  obvious  to  any 
thoughtful  mind,  and  are  very  plain,  such  as  sufficient  supply  of 
ore  at  a  fixed  and  low  rate  for  a  reasonable  length  of  time,  and 
contiguity  to  water,  railroads,  or  some  cheap  means  of  transporta- 
tion to  the  manufactory,  which,  owing  to  the  nature  of  its  product, 
must  be  situated  in  the  immediate  vicinity  of  its  market. 

The  technical  conditions,  though  more  numerous,  are  almost 
equally  easy  of  comprehension.  An  almost  absolute  freedom  from 
gangue  is  essential,  for  the  simple  reason  that  the  presence  of  for- 
eign substances  lowers  the  percentage  of  sulphur  and  necessitates 
the  handling  of  worthless  material,  thus  lessening  the  capacity  of 
the  works  and  producing  other  unfavorable  results.  For  the  same 
reason,  though  in  a  less  degree,  the  presence  of  any  other  sulphides 
but  the  bisulphide  of  iron,  which  forms  the  ore  proper,  is  disad- 
vantageous; for  no  other  compound  of  sulphur  contains  either  so 
high  a  percentage  of  the  same  or  parts  with  it  so  freely.  Even 
the  copper  pyrites,  which  in  many  instances  forms  the  principal 
value  of  the  ore,  is  detrimental  to  the  manufacture  of  sulphuric 
acid,  both  because  it  contains  less  sulphur  and  because  it  is  too 
fusible  to  permit  the  proper  regulation  of  the  temperature.  Be- 
yond the  limit  of  8  per  cent,  of  copper  in  the  pyrites,  it  cannot  be 
profitably  employed  in  the  manufacture  of  acid.  The  Spanish 
pyrites,  from  which  so  large  a  proportion  of  the  acid  produced  in 
England  is  made,  contains  on  an  average  about  3  per  cent,  of 
copper,  and  about  48  per  cent,  of  sulphur,  this  remarkably  high 
percentage  of  sulphur  showing  its  freedom  from  gangue. 

An  analysis  of  the  average  ore  from  the  celebrated  Rio  Tinto 
mine  may  be  of  interest,  as  a  type  of  a  very  favorable  cupriferous 
pyrite  for  acid  making: 



Sulphur 48.00 

Iron  40.74 

Copper 3.42 

Lead 0.82 

Lime 0.21  !  

Total 100.15 

Magnesia 0.08 

Arsenic 0.21 

Insoluble 5.67 

Oxygen  and  moisture 1.00 

The  ore  used  by  several  large  acid-works  in  Boston  and  New 
York  is  obtained  principally  from  Canada,  some  thirty  miles  from 
the  Vermont  line,  and  although  somewhat  variable  in  purity, 
averages  about  3.5  per  cent,  of  copper  and  45  per  cent,  of  sulphur, 
the  percentage  of  gangue  being  greater  than  m  the  Spanish  ores. 

An  excellent  quality  of  pyrites  is  mined  from  a  large  deposit  in 
Western  Massachusetts,  and  in  both  Virginia  and  Georgia  are  beds 
of  pyrites  now  under  process  of  development,  which,  on  competent 
authority,  are  said  to  rival  the  Spanish  mines  in  quality. 

The  presence  of  arsenic  and  antimony  has  a  deleterious  effect  on 
the  quality  of  the  resulting  acid,  while  lead  heightens  the  fusibility 
of  the  charge,  besides  wasting  sulphur  by  forming  a  stable  lead 
sulphate,  and  any  foreign  substance,  however  harmless  otherwise, 
lessens  the  percentage  of  sulphur. 

An  important  point,  sometimes  overlooked  by  non-professionals 
in  determining  the  value  of  a  sample  of  pyrites,  is  its  mechanical 
behavior  during  the  process  both  of  crushing  and  of  roasting.  A 
granular  ore,  soft  or  easily  disintegrated,  will  increase  the  propor- 
tion of  fines,  which,  altliough  now  utilized  with  great  success  in 
the  manufacture  of  acid,  are  still  undesirable  as  requiring  a  more 
expensive  plant  and  entailing  a  greater  cost  in  their  treatment. 
A  still  more  serious  production  of  fines  may  take  place  in  the  kiln 
itself  in  the  case  of  ores  that  decrepitate,  sometimes  occurring  to 
such  an  extent  as  entirely  to  choke  the  draught  and  render  their 
emplovinent  impossible. 

One  of  the  most  serious  errors  ever  perpetrated  in  the  manufacture 
of  acid  from  pyrites  is  the  attempted  employment  of  pyrrhotite, 
or  monosulphide  of  iron,  for  pyrite — bisuliihide  of  iron.  Aside 
from  the  greatly  lessened  proportion  of  sulphur,  3G  per  cent,  as 
against  53  per  cent.,  the  monosulphide  will  not  even  yield  freely 
what  sulphur  it  contains,  but  crusts  with  oxide  of  iron,  tnrns 
black,  and  is  soon  extinguished  when  treated  in  an  ordinary  pyrites 
kiln.     It  seems  scarcely  possible  that  extensive  works  for  the  man- 

THE    ROASTINU    OF   ORES   IN    LUMP   FORM.  169 

ofacture  of  sulphuric  acid  (and  copper)  should  have  been  erected, 
their  ore  supply  being  entirely  derived  from  a  deposit  of  the  value- 
less mouosulphide;  but  such  has  been  the  case  in  more  than  one 
instance,  and  will  continue  to  be  so  in  enterprises  conducted  with- 
out the  aid  of  skilled  direction.  One  of  the  most  striking  instances 
of  this  kind  is  a  now  extinct  Massachusetts  company,  which  is  said 
to  have  expended  over  $200, OOU  in  this  manner,  all  of  which  was  a 
total  loss,  excepting  the  small  amount  realized  from  the  sale  of 
buildings  and  land. 

Under  certain  conditions,  the  use  of  kilns  for  the  calcination  of 
cupriferous  pyrites  without  the  production  of  sulphuric  acid  may 
be  found  advantageous,  as  in  the  case  of  the  former  Orford  Nickel 
and  Copper  Company,  near  Sherbrooke,  Province  of  Quebec,  which, 
after  employing  heap-roasting  for  some  time,  erected  a  large  num- 
ber of  kilns  solely  for  the  purpose  of  calcining  its  ore  previous  to 
smelting;  finding  the  saving  in  time  and  avoidance  of  waste,  com- 
bined with  the  lessening  of  the  annoyance  formerly  experienced 
from  sulphur  fumes,  a  sufficient  advantage  to  repay  the  somewhat 
heavy  cost  of  the  burners. 

The  minimum  percentage  of  sulphur  sufficient  to  maintain 
combustion  in  kilns  does  not  yet  seem  to  have  been  positively  de- 
termined; but  with  an  ore  otherwise  favorable,  it  is  probable  that 
25  per  cent,  is  quite  sufficient  for  the  purpose. 

For  economy's  sake,  as  well  as  for  the  purpose  of  retaining  the 
heat,  kilns  are  constructed  in  blocks  of  considerable  length  and  of 
the  depth  of  two  burners,  the  front  of  each  facing  outward,  while 
the  flue  in  which  the  gas  is  conveyed  to  its  destination  is  built  on 
top  of  the  longitudinal  center  wall.  Fire-brick  are  used  wherever 
the  masonry  is  exposed  to  heat  or  wear,  and  the  entire  block  of 
furnaces  is  surrounded  by  cast-iron  plates,  firmly  bolted  in  posi- 
tion, and  provided  with  the  necessary  openings  for  manipulation. 

No  fuel  is  required  after  the  burners  are  once  in  operation;  and 
when  in  normal  condition,  the  attendance  demanded,  aside  from 
the  labor  connected  with  the  regular  charge  of  from  500  to  2,000 
pounds  of  ore  once  in  twelve  or  twenty-four  hours,  is  very  slight. 

Much  skill  and  experience,  however,  are  required  to  maintain 
the  regular  working  of  the  kilns,  especially  with  ores  that  are  not 
exactly  suited  to  the  process. 

From  5  to  10  per  cent,  of  fines  may  also  be  desulphurized  with 
the  coarse  ore  without  seriously  interfering  with  the  process. 
They  are  thrown  toward  the  back  and  sides  of  the  shaft,  leaving 


the  center  uncovered;  otherwise,  the  draught  is  affected  and 
serious  irregularities  supervene. 

In  accordance  with  the  policy  adopted  throughout  this  work, 
no  detailed  estimate  of  expense  will  be  given  in  the  few  instances 
where  the  author  is  unable  to  base  the  same  on  pei-sonal  experience. 

Such  is  the  case  in  kiln  roasting;  but  we  are  assured  by  the  best 
authorities  that  the  expense  of  calcination  by  this  method  does  not 
exceed  that  of  stall  roasting,  though  the  first  cost  of  the  plant  is 
considerably  greater. 

The  results  obtained  by  this  process  are  unexampled  in  the 
roasting  of  lump  ores,  although  there  is  no  doubt  that  a  consider- 
able share  of  the  success  is  due  to  the  fact  that  the  sulphur  is  the 
object  of  interest,  instead  of  merely  being  a  waste  product  to  be 
driven  off  as  far  as  convenient. 

If  more  than  4  per  cent,  of  sulphur  remains  in  the  cinders,  as 
the  residue  from  this  process  is  called,  the  result  is  not  considered 
satisfactory.  It  is  needless  to  say  that  such  a  perfect  desulphuri- 
zation  cannot  be  obtained  in  either  heap  or  stall-roasting,  nor  is  it 
necessary  or,  in  many  cases,  even  beneficial  for  the  subsequent 
process,  although,  of  course,  in  most  instances  the  lack  of  sulphur 
in  the  furnace  charge  forms  a  welcome  outlet  for  Jthe  admixture  of 
raw  fines,  which  may  thus  escape  the  expense  of  calcination. 

Within  the  past  few  years,  the  utilization  of  these  fines  has  at- 
tracted much  attention,  and  the  efforts  to  calcine  them  in  automatic 
furnaces  for  the  production  of  sulphurous  acid  have  been  crowned 
with  success,  as  will  be  again  alluded  to  when  treating  of  the 
"Roasting  of  Pulverized  Materials." 

The  attempt  to  utilize  kilns,  with  certain  slight  modifications, 
for  the  roasting  of  copper  matte  has,  after  many  difficulties  and 
much  expense,  attained  a  successful  issue  at  certain  European 
works,  especially  at  the  Mansfeld  copper  works  in  Germany,  the 
object  in  view  being  rather  the  abolition  of  the  nuisance  arising 
from  the  escape  of  the  sulphur  fumes  into  the  atmosphere  than 
anv  expectation  of  financial  advantage  from  the  employment  of  a 
substance  so  poor  in  sulphur  for  the  manufacture  of  acid. 



At  the  beginning  of  Chapter  V.  we  classified  the  apparatus 
suitable  for  roasting  ores  in  a  finely  divided  form,  as  follows: 

1.  Shaft-furnaces. 

2.  Stalls. 

3.  Hand  reverberatory  calciners. 

{a)  Open  hearth. 
{b)  Muffle. 

4.  Revolving  cylinders. 

(a)  Continuous  discharge. 

(b)  luterniitteut  discharge. 

5.  Automatic  reverberatory  calciners. 

(«)  Stationary  hearth. 
{b)  Revolving  hearth. 


This  group  includes  some  of  the  most  important  and  useful 
appliances  for  the  roasting  of  sulphureted  substances,  where  the 
utilization  of  the  fumes  for  the  manufacture  of  sulphuric  acid 
forms  a  part  of  the  process  of  calcination. 

If  the  question  of  acid  manufacture  be  left  entirely  out  of  con- 
sideration, and  the  comparative  economy  of  each  method  of  calci- 
nation be  judged  solely  upon  its  own  merits,  it  is  doubtful  v/hether 
resort  would  be  had  to  these  furnaces,  save  under  exceptional  con- 
ditions, as  their  limited  capacity,  great  cost  of  construction,  and 
imperfect  work,  except  under  the  most  skillful  managemciit, 
would  effectually  bar  their  introduction.  But  under  the  stimulus 
arising  from  the  enforced  manufacture  of  acid  from  pulverized 
pyrites,  and  the  consequent  necessity  of  employing  some  form  of 
automatic  furnace  in  which  the  gases  arising  from  the  oxidation  of 
the  ore  are  kept  separate  from  the  products  of  combustion  of  the 
fuel,  this  type  of  calciner  has  received  such  attention  and  study 
that  it  pro;nises  fairly  to  rival  the  most  economical  form  of  roast- 


iug  apparatus  knowu  to  metallurgy.  Tlie  studeut  is  referred  to 
Lunge's  work  on  the  manufacture  of  sulphuric  acid  for  full  details 
regarding  this  and  other  forms  of  furnace  suited  to  the  calcination 
•of  ores  in  connection  with  acid-making. 

TJie  Gerstenhofer  shelf -furnace  was  the  first  successful  calciner 
of  this  type,  and  is  still  largely  nsed,  though  becoming  gradually 
supplanted  by  improved  modifications.  The  few  furnaces  of  this 
pattern  that  have  been  constructed  in  the  United  States  have 
failed  to  answer  the  desired  purjaose,  owing  to  imperfect  construc- 
tion, poor  refractory  materials,  and  want  of  skill  in  management. 
The  Gerstenhofer  furnace  consists  of  a  vertical  shaft,  surmounted 
by  a  mechanical  device  for  feeding  the  pulverized  sulphides  in  any 
desired  quantity,  ar.d  contains  a  great  number  of  parallel  clay 
leds'es  of  a  triangular  form,  one  of  the  flat  surfaces  being  placed 
uppermost.  These  are  so  arranged  as  to  obstruct  the  ore  in  its 
passage,  and  delay  it  sufficiently  to  effect  a  certain  degree  of  oxida- 
tion, which  is  seldom  perfect  enough  to  yield  the  desired  result 
without  a  supplementary  calcination  in  some  other  form  of  fur- 
nace. The  front  wall  of  the  shaft  is  pierced  by  parallel  rows  of 
rectangular  openings,  for  the  purpose  of  changing  the  clay  sb.elves 
or  of  cleansing  the  same. 

The  oxidation  of  the  sulphides  generates  sufficient  heat  for  the 
proper  working  of  the  process,  so  that  the  sulphurous  gases  may 
be  obtained  for  the  manufacture  of  acid  free  from  any  products  of 
the  combustion  of  fuel. 

Though  greater  capacity  has  been  reported,  I  have  never  been 
able  to  satisfy  myself  that  a  full-sized  Gerstenhofer  could  handle 
more  than  6,000  pounds  per  24  hours  of  granular  pyrites,  with  46 
per  cent,  sulphur,  the  residues  averaging  6  per  cent,  sulphur. 

Hasenclever'' s  fiirnace  consists  of  a  vertical  shaft  containing  six 
to  eight  inclined  fire-clay  shelves  arranged  in  a  zigzag  fashion,  as 
in  Stetefeldt's  dry  kiln.  The  angle  of  inclination  is  about  40  de- 
grees, so  that  a  thin  layer  of  ore  covers  each  shelf,  descending  by 
its  own  gravity  as  rapidly  as  the  finished  product  is  carried  out  at 
the  bottom  by  a  fluted  roller.  The  discharge  is  continuous,  but 
the  capacity  very  limited,  seldom  reaching  one  ton  of  fines  per  24 
hours.  Nor  can  the  heat  be  maintained  without  extraneous  aid, 
which  is  usually  supplied  by  connecting  it  with  a  kiln  in  which 
lump  pyrites  is  burned. 

Tlie  Maletrafunutre  properly  belongs  under  "Hand  Reverbera- 
tories,"  as  it  consists  of  a  number  of  snuiU  hearths  one  above  the 


other,  over  which  the  ore  is  moved  by  hand-rakes.  This  furnace 
is  interesting  as  roasting  fines  entirely  without  the  aid  of  extrane- 
ous heat.  It  will  roast  about  one  ton  of  heavy  pyrites  fines  per 
24  hours  down  to  3  per  cent,  sulphur, 

77ie  Stetefeldt  furnnrj,  so  invaluable  for  the  chloridizing-roasting 
of  silver  ores,  is  a  shaft  provided  with  a  grate  for  the  generation 
of  such  a  degree  of  temperature  as  would  be  lacking  in  the  roasting 
of  ores  so  poor  in  sulphur  as  those  usually  exposed  to  this  treat- 
ment, as  well  as  an  auxiliary  fireplace  for  the  more  perfect  chlori- 
diziiig  of  the  flue-dust,  which,  owing  to  the  fine  pulverization  of 
the  ore  and  the  strong  draught  essential  to  the  proper  working  of 
the  apparatus,  is  formed  in  unprecedented  amounts,  and  pretty 
thorouglily  regained  in  ample  dust-chambers. 

The  employment  of  an  auxiliary  fireplace,  and  the  invention  of 
a  highly  ingenious  and  perfect  automatic  ore-feeder,  constitute 
important  claims  to  originality  that  are  frequently  overlooked  by 
writers  in  commenting  on  tliis  furnace.  Its  capacity  is  very  great, 
60  tons  in  '2.Al  hours  being  easily  worked  in  one  of  the  large-sized 
furnaces  of  this  type;  and  were  it  possible  to  obtain  equally  good 
results  by  employing  it  for  oxidizing-roasting,  it  would  be  tlie 
most  valuable  addition  to  the  modern  metallurgy  of  copper.  But 
as  it  is  at  the  present  time,  it  cannot  be  enumerated  among  the 
resources  of  the  copper  smelter,  although  late  experiments  indicate 
the  probability  of  its  snccessful  adaptation  to  this  jjurpose. 


Pelatan's  roasting  and  agglomeration  furnace  presents  some 
novel  features,  and  is  intended  for  the  calcination  of  pyritic 
smalls  or  other  sulphides.  It  consists  of  a  long,  narrow  covered 
stall,  provided  with  sheet-iron  front,  and  cast-iron  side-doors.  It 
contains  a  close-meshed  grate.  The  charge  of  fine  or  granular  oro, 
after  being  placed  in  the  grate,  is  ignited  from  below.  A  light 
blast  is  used  under  the  grate,  and  it  is  claimed  that  while  the 
plant  is  cheap,  the  capacity  is  large,  and  that  a  10-ton  charge  of 
most  ores  can  be  well  roasted  and  slightly  agglomerated  in  12 
hours.  If  such  results  could  be  constantly  obtained  in  practice, 
the  apparatus  would  be  of  much  value  in  many  places.  Good  re- 
sults are  reported  from  the  Laurium  galena  mines  in  Greece,  and 
from  pyrites  mines  in  Chili,  that  are  working  up  their  old  piles  of 
fines  in  this  furnace. 



{n)  With  Open  Hearth.  —  The  esseutial  features  of  the  ordinary 
reverberatory  calciuer  are  a  hearth,  heated  by  a  lireplace,  from 
which  it  is  ordinarily  separated  by  the  bridge-wall,  and  accessible 
by  certain  openings  through  the  side  walls,  the  whole  being  covered 
by  a  flat  arch  against  which  the  flame  revei'herates  in  its  passage 
from  the  grate  to  the  flue,  thus  being  brought  momentarily  in 
contact  with  the  ore  spread  upon  the  hearth,  while  the  combined 
gases  from  fuel  and  charge  pass  into  the  open  air  through  a  chim- 
ney, in  many  cases  first  traversing  a  series  of  flues  and  chambers 
for  the  purpose  of  retaining  such  particles  of  metal  as  may  have 
been  either  chemically  or  mechanically  borne  away  by  the  rapid 

A  very  small  grate  surface,  as  compared  with  the  hearth  area, 
distinguishes  this  type  from  the  reverberatory  smelting-furnacf, 
and  corresponds  to  the  very  moderate  temperature  suited  to  the 
process  of  calcination,  permitting  its  almost  entire  construction  of 
common  red  brick. 

A  single  detailed  account  of  the  longest  and  largest  variety  of 
calciner  in  common  use  will  serve  as  a  model  for  all  smaller  speci- 
mens of  the  same  class. 

The  principal  variable  dimension  of  a  copper-desulphurizing 
furnace  is  its  length,  as,  for  economical  reasons,  its  width  should 
always  be  as  great  as  is  compatible  with  convenient  manipulation. 
Experience  has  placed  this  limit  at  16  feet  for  the  inside  measure- 
ment of  the  hearth,  nor  should  this  dimension  be  lessened  without 
good  and  sufficient  reasons. 

The  length  of  the  hearth  is  limited  chiefly  by  the  capacity  of 
the  ore  to  generate  heat  during  its  oxidation,  the  immediate  influ- 
ence of  the  fireplace  being  seldom  capable  of  maintaining  the 
requisite  temperature  upon  a  hearth  over  35  feet  in  length, 
without  resorting  to  the  use  of  a  forced  blast,  or  of  a  draught  so 
powerful  as  greatly  to  increase  the  loss  in  dust  as  well  as  the  con- 
sumption of  fuel. 

The  importance  of  the  heat  generated  by  the  oxidation  of  sul- 
phides in  maintaining  a  proper  temperature,  and  especially  in  con- 
veying the  heat  to  a  great  distance  from  the  initial  point,  is  not 
always  appreciated.  Its  intensity  and  durability  depend  upon  the 
percentage  of  sulphur  in  the  ore,  and  also  not  a  little  upon  the 
manner  in  which  it  is  chemically  combined,  the  bisulphides — suc^ 




as  iron  pyrites — furuishing  a  much  greater  amount  of  heat  thap 
mouosulphides  coutaiuiug  an  equal  gross  amount  of  sulphur. 

An  ore  carrying  less  than  10  per  cent,  of  sulphur  will  not  furnish 
sufficient  heat  to  warrant  the  addition  of  a  second  hearth  to  the  first 
16  feet,  wbitih  will  be  assumed  as  the  normal  length  of  a  single 
hearth.  (Such  a  condition  would  scarcely  occur  in  practice,  as, 
under  ordinary  circumstances,  any  copper  ore  containing  snch  a 
low  percentage  of  sulphur  could  be  smelted  raw.)  An  increase  of 
sulphur  to  15  per  cent.,  however,  will  be  sufficient  to  heat  the 
second  hearth,  while  a  20  per  cent,  sulphur  ore  should  work  rapidly 
in  a  three-hearth  furnace.  The  addition  of  a  fourth  and  final 
section  is  rendered  justifiable  by  the  increase  of  the  average  sul- 
phur contents  of  the  ore  to  25  per  cent.,  and  even  a  20  per  cent, 
bisulphide  charge  may  be  worked  to  advantage  in  the  same. 

The  adoption  of  this  method  of  roasting,  by  which  the  ore  is  fed 
into  one  end  of  the  furnace,  and  gradually  moved  to  the  other 
extremity  before  discharging,  is  attended  with  several  obvious 
advantages,  among  which  are:  The  gradual  elevation  of  tempera- 
ture from  a  point  compatible  with  the  easy  fusibility  of  the  unal- 
tered suphides  to  that  degree  necessary  for  the  complete  decompo- 
sition of  the  pertinacious  basic  sulphates  of  copper  and  zinc;  the 
great  saving  in  fuel  effected  by  thus  obtaining  the  full  benefit  of  the 
heat  generated  in  the  process  of  roasting  itself;  the  certainty  that 
thechargemust  undergo  a  certain  number  of  thorough  stirrings  and 
turnings  in  its  transportation  over  so  extended  a  space;  the  estab- 
lishment of  a  fixed  duty,  which  must  be  performed  by  the  work- 
men, whose  labor  can  thus  be  much  more  easily  controlled  than 
with  the  single-hearthed  type  of  calciner,  where  the  attendants 
can  easily  substitute  an  idle  scratching  for  the  vigorous  manipula- 
tion necessary  to  move  the  ore  forward  promptly;  a  great  simpli- 
fication in  firing,  it  being  only  necessary  in  the  long  furnace  to 
maintain  an  even,  high  temperature,  while  with  the  single  hearth, 
much  experience  and  judgment  are  required  to  adapt  the  heat  to 
the  ever-varing  condition  of  the  charge;  lastly,  a  decided  economy 
in  construction,  the  ratio  of  fire-brick  to  common  red  brick  for  an 
equal  capacity  of  plant  being  much  less  in  the  employment  of  long 

As  there  seems  to  be  almost  no  limit  to  the  extent  of  surface 
over  which  the  requisite  temperature  may  be  obtained  in  the  calci- 
nation of  highly  sulphureted  ores,  it  is  very  natural  that  experi- 
ments shoujd  have  been  made  with  still  longer  furnaces  than  any 


yet  mentioued,  120  feet  being  the  extreme  inside  leugtli  yet  at- 
tempted, so  far  as  known  to  the  writer;  but  careful  and  repeated 
trials  have  shown  beyond  a  doubt  that  no  sufficient  advantage  is 
reaped  to  pay  the  increased  cost  of  the  enclosing  building  and  other 
expenses  of  plant.  It  is  not  possible  for  two  attendants  properly 
to  manage  a  furnace  having  more  than  four  full-sized  hearths,  if 
the  latter  is  pushed  to  its  full  capacity,  while  the  addition  of  a 
fifth  hearth  demands  a  third  laborer,  whose  time,  however,  will 
not  be  fully  occupied,  while  a  sixth  hearth  will  overtax  the  three 
workmen.  In  short,  the  testimony  of  many  excellent  metallur- 
gists, to  which  the  author  can  add  his  own  exp  rience,  unequivo- 
cally condemns  the  lengthening  of  ordinary  calciniiig-furiuices 
beypnd  the  limits  above  indicated,  excepting  under  special  and 
peculiar  conditions. 

The  number  of  working-doors  to  a  long  calcining-furnace,  where 
the  ore  is  moved  from  rear  to  front,  should  be  as  few  as  possihle. 
The  limit  for  comfortable  work  should  not  exceed  8  feet  between 
centers  of  doors,  and  any  distance  less  than  6  feet  is  a  decided 

The  sides  of  the  working-door  frames  should  have  short  lugs, 
not  exceeding  6  inches  in  length,  cast  on  them,  in  order  that  they 
may  be  firmly  held  in  position  by  the  buckstaves,  which  are  placed 
in  pairs  for  this  purpose,  a  single  buckstaff  being  placed  in  the 
center  of  the  space  between  each  pair.  The  bottom  of  the  door- 
frames should  be  on  a  level  with  the  hearth  surface,  which  should 
be  3  feet  above  the  floor  grade  of  the  building,  and  should  slope 
gradually  upward  toward  the  rear  of  the  furnace,  to  correspond 
with  the  increased  height  of  each  succeeding  hearth. 

The  common  practice  of  filling  up  the  portions  of  the  hearth 
between  the  working-doors  with  projecting,  triangular  masses  of 
brick-work  cannot  be  recommended,  as  valuable  space  is  often 
sacrificed  in  this  manner.  Slight  projections,  as  shown  in  the 
cut,  may  be  built  to  fill  the  absolutely  inaccessible  angles; 
but  with  properly  constructed  door-frames,  and  careful  manipula- 
tion on  the  part  of  the  roasting  attendants,  but  little  waste  area 
should  exist,  and  this  will  regulate  itself  by  becoming  filled  with 
ore,  which  may  remain  there  permanently.  This  refers,  of  course, 
to  the  treatment  of  large  quantities  of  low-grade  ores,  where  slight 
inaccuracies  resulting  from  the  trifling  mixture  of  ores  can  do  no 

*  Tbes-?  building  directions  are,  in  the  main,  equally  applicable  to  any  of 
the  automatic  calcining:  furnaces. 


After  raising  the  side  walls  to  the  height  reqAiired  by  the  iron 
door-frames,  usually  about  ten  inches  above  the  hearth  level,  th.e 
skewback  for  the  main  arch  should  be  laid.  This  applies  to  the 
entire  furnace  from  the  beginning  of  the  fire-box  to  the  extremity 
of  the  rear  hearth,  and  is  a  very  simple  matter,  especially  if  the 
arch  is  to  be  perfectly  horizontal,  as  is  to  be  recommended  in  n^o.^t 
cases.  A  taut  line  should  be  stretched,  to  insure  accurate  work, 
and  if  red  brick  are  used,  they  should  be  cut  on  one  long  edge, 
being  laid,  of  course,  longitudinally  and  on  the  flat.  Tiiey  should 
be  cut  at  an  angle  slightly  greater  than  required  by  the  curve  of 
the  arch,  which  should  rise  about  three-quarters  of  an  inch  to  the 
foot,  making  a  IG-foot  arch  12  inches  higher  in  the  center  than  at 
the  sides.  This  rise,  though  less  than  is  often  recommended,  will 
be  found  ample  to  insure  perfect  safety  and  durability,  and  will 
tend  to  spread  the  flame  and  heat  toward  the  sides  of  the  hearth. 

If  so-called  "side  skewback"  flre-brick  are  within  reach,  they 
should  be  used  in  place  of  the  red  brick,  saving  much  cutting  and 
insuring  a  better  job.  Three  rows,  in  height,  of  red  brick,  or  two 
of  fire-brick,  will  give  a  solid  bearing,  the  total  number  required 
for  a  furnace  of  the  size  under  consideration  being,  respectively, 
600  and  375. 

It  is  of  sufficient  importance  to  bear  repetition,  that  all  portions 
of  the  mason  work  above  the  hearth  line,  or  wherever  exposed  to 
heat,  must  be  laid  in  clay — common  brick  clay,  tempered  with 
sand,  being  quite  good  enough  for  all  portions  of  the  furnace — as 
fire-clay  is  usually  expensive  in  the  localities  where  copper  ores 

Lime  mortar,  much  improved  by  the  admixture  of  a  little  good 
cement — say  10  per  cent. — may  be  advantageously  employed  for 
the  outside  work  and  wherever  there  is  no  danger  of  heat,  as  it 
makes  handsomer  and  stronger  work,  and  is  greatly  preferred  by 
the  masons,  who  require  constant  supervision  to  compel  them  to 
use  clay  mortar  where  it  is  necessary. 

The  heavv  iron  bridge-plate,  so  indispensable  in  reverberatory 
smelting-furnnces,  may  be  entirely  omitted,  the  bridge  being  built 
up  solid  and  covered  on  tlie  top  and  sides  with  fire-brick,  with  the 
exception  of  a  longitudinal  opening  3  by  8  inches,  which  should 
penetrate  it  from  one  end  to  the  other,  communicating  with  the 
outside  air  on  each  side  of  the  furnace,  and  with  the  hearth  by  some 
ilnzen  2  by  4-inch  openings. 

Rv  this  means,  Vented  air  free  from  all  reducing  gases  is  admitted 


into  tbfci  furnace  below  the  sheet  of  flame  that  sweeps  over  the  top 
of  the  bridge.  The  oxidizing  effect  of  this  current  of  air  is  very 
powerful,  and,  as  frequently  determined  by  experiment,  hastens 
materially  the  calcining  process. 

If  wood  is  used  as  a  fuel,  an  additional  row  of  similar  openings 
should  be  constructed  in  the  arch,  immediately  over  the  front  line 
■of  the  bridge- wall,  by  which  a  much  more  perfect  combastion  of 
the  gases  is  effected.  With  coal  as  a  fuel,  the  latter  openings  are 
less  urgent,  provided  the  firing  is  properly  managed. 

Aside  from  the  16  working-door  frames,  and  the  ordinary  doors 
for  fire-box  and  ash-pit,  no  castings  are  necessary  for  the  entire 
structure,  excepting  a  small  frame  to  protect  the  charging-hole, 
which  should  be  situated  a  little  back  of  the  center  of  the  rear 
hearth,  being  placed,  of  course,  in  the  medium  longitudinal  line 
■of  the  furnace.  It  will  add  also  materially  to  the  durability  of 
the  fire-box  to  surround  the  portions  of  the  same  most  exposed  to 
pressure  or  mechanical  violence,  by  light  cast  plates,  held  in  place 
by  the  uprights. 

As  the  portion  of  the  hearth  immediately  beneath  the  charging- 
hole  is  exposed  to  excessive  wear  from  the  constant  precipitation 
of  heavy  masses  of  wet  material  upon  it,  an  area  some  six  feet 
square,  and  in  the  locality  designated,  should  be  constructed  of 
either  fire-brick  or  slag  blocks,  the  latter,  from  their  texture  and 
general  physical  condition,  being  peculiarly  well  suited  to  the 

By  referring  to  Fig.  23,  it  can  be  plainly  seen  at  what 
stage  in  construction  the  various  bearing- bars  and  other  iron 
work  must  be  inserted. 

It  will  be  noticed  that,  instead  of  adopting  the  ordinary  large 
ash-pit,  entirely  open  at  the  rear,  according  to  the  invariable  Eng- 
lish practice^  preference  is  given  to  a  closed  ash-pit,  to  which  air 
is  admitted  by  a  door  at  one  or  both  ends.  This  effects  a  great 
saving  in  fuel,  and  brings  the  process  of  combustion  more  perfectly 
under  control.  Comparative  tests,  extending  over  a  considerable 
period  of  time,  show  this  saving  to  amount  to  about  50  per  cent, 
of  the  entire  fuel  consumed,  in  the  case  of  coal,  and  about  65  per 
cent,  (in  volume)  when  pine  wood  is  used.  The  tight  ash-pit 
becomes,  of  course,  a  matter  of  positive  necessity  where  anthracite 
coal,  with  a  forced  blast,  is  used. 

The  side  and  end  walls  having  been  carried  up  to  the  required 
lieight,  and  the  skewback  constructed  on  both  sides  for  their  entire 


length,  the  carpenters  take  possession  temporarily,  nsually  under 
the  supervision  of  the  head  mason,  to  put  in  the  wooden  center  on 
which  the  arch  is  to  be  built.  If  a  second  furnace,  or  indeed  any 
other  work,  is  available  for  the  remaining  masons,  it  is  advanta- 
geous, though  not  indispensable,  to  permit  the  furnace  to  stand 
uncovered  for  several  days,  thus  allowing  the  mortar  to  set,  and 
greatly  increasing  the  strength  of  the  mason  work. 

Having  selected  for  description  that  pattern  of  calciuer  in  which 
the  gradual  diminution  of  the  space  between  arch  and  hearth,  as 
it  recedes  from  the  grate,  is  due  to  successive  slight  elevations  of 
the  hearth  level,  instead  of  the  ordinary  downward  pitch  of  the 
roof,  it  is  evident  that  the  arch  throughout  its  entire  extent  will 
be  horizontal,  while  all  four  inclosing  walls  are  built  up  to  the 
same  height  at  every  point,  with  the  exception  of  a  rectangular 
flue-opeuing  in  the  rear  wall,  G  by  30  inches. 

The  construction  of  the  wooden  pattern  or  center  is,  therefore^ 
extremely  simple,  requiring  only  some  20  pieces  of  2-inch  plank, 
16  feet  long  and  14  inches  wide;  a  lot  of  2  by  4  scautling,  to  form 
posts  about  10  inches  in  length,  four  of  these  being  needed  to  sup- 
port each  plank  on  edge;  and  finally,  a  sufficient  amount  of  4:-inch 
battens,  from  one-half  to  one  inch  thick,  to  cover  the  area  of  the 
required  roof,  when  placed  about  three-quarters  of  an  inch  apart. 
The  planks  should  be  perfectly  sound,  and  at  least  partially 

By  the  aid  of  a  long  rod,  moving  upon  a  pivot  at  one  end,  while 
the  free  extremity  carries  a  pencil,  a  segment  of  a  circle  corre- 
sponding to  a  rise  of  I'-i  inches  in  the  center  of  the  length  of  10 
feet,  should  be  struck  on  each  plank,  and  the  line  followed  accu- 
rately with  a  jig-saw. 

The  segments  for  that  portion  of  the  arch  over  the  bridge  and 
lire-box  are  shorter,  of  course,  than  those  belonging  to  the  main 
hearth,  but  should  be  got  out  in  the  same  manner,  and  then  shoit- 
enetl  at  each  end  to  the  required  length. 

The  scantling  should  be  cut  into  posts  somewhat  shorter  than 
necessary  to  bring  the  curve  on  the  upper  edge  of  the  segments  to 
the  proper  height  for  the  lower  surface  of  the  arch,  so  that  each 
post  may  be  wedged  to  an  exact  bearing  with  thin  slips  of  wood. 
It  is  quite  necessary  that  the  weight  should  be  evenly  distributed, 
and  each  segment,  when  brought  into  correct  position,  is  held 
there  by  driving  a  nail  through  a  longitudinal  line  of  battens  in 
the  center  and  at  each  extremity. 


The  segiiieats  for  slopiug  arches  sliould  be  still  further  strengtli- 
eued  by  short  braces  toe-nailetl  obliquely  from  the  upper  etlge  of 
one  strip  to  the  lower  edge  of  its  neighbor,  aud  so  on  throughout 
the  entire  frame. 

An  omission  of  this  precaution  once  caused  the  canting  of  the 
segments  and  consequent  destruction  of  a  large,  nearly  completed 
arch  under  the  author's  charge. 

No  diflBculty  will  be  experienced  in  removing  the  wooden  pattern 
in  good  condition  for  farther  use,  provided  it  is  supported  on  small 
posts  as  just  described;  but  if  long,  heavy  blocks  of  timber  are 
used  as  a  foundation  for  the  segments,  great  labor  as  well  as  much 
injurious  sledging  must  accompany  their  removal,  resulting  usually 
in  the  comjilete  destruction  of  both  segments  and  battens.  In 
fact,  where  this  method  of  support  has  been  practised,  it  will  be 
found  best  to  burn  out  the  enclosed  patterns,  after  the  tie-rods  are 
properly  tightened,  closing  both  damper  and  ash-pit  so  as  to  allow 
only  a  slow  smoldering,  and  prevent  any  injurious  rise  of  tempera- 
ture in  the  still  damp  furnace. 

Few  jobs  of  mason  work  require  more  care  and  conscientiousness 
than  the  laying  of  a  large  calciner  arch,  as,  owing  to  its  great 
width  and  slight  curvature,  a  very  little  lack  of  closeness  in  its 
myriad  joints  would  be  sufficient  to  allow  it  to  yield  to  the  enor- 
mous pressure  brought  to  bear  by  its  own  weight,  and  become 
sutiiciently  compressed  to  slip  down  between  its  side  walls.  It  is 
quite  a  simple  matter  to  lay  a  good  solid  arch  of  fire-brick,  owing 
to  their  great  regularity  aud  smoothness,  and  almost  perfect  rectan- 
gular form;  but  when  red  brick  are  used,  which  vary  so  in  size 
and  thickness,  and  are  so  frequently  warped  out  of  all  reasonable 
shape,  much  care  is  required. 

In  ordinary  calciners,  it  is  customary  to  construct  that  portion 
of  the  arch  from  the  fire  end  of  the  furnace  to  a  point  midway 
between  the  first  and  second  working-doors  of  fire-brick,  nine 
inches  in  thickness,  the  brick  standing  endwise.  At  this  point, 
or  even  considerably  sooner,  when  necessary,  red  brick  are  substi- 
tuted, being  placed  also  on  end,  each  brick,  after  being  dip})ed 
into  a  pail  of  liquid  clay  mortar,  being  pressed  closely  against  its 
neighbor,  and  finally  settled  into  position  with  a  few  light  blows 
of  the  hammer. 

Moderately  soft  brick  are,  as  a  rule,  best  suited  to  this  purpose, 
although  they  must,  of  course,  possess  ample  solidity  to  resist  the 
compression    to    which    they    are   exposed.      Hard-burned    brick, 


though  stronger,  are  too  irregular  and  AViuped  to  be  often  used  in 
a  large  arch,  and  in  auv  case  the  brick  should  be  all  carefull}'  selected 
beforehand  by  the  attendant,  and  assorted  in  sucli  a  manner  that 
each  longitudinal  row — extending  the  entire  length  of  the  furnace 
— is  composed  of  brick  of  about  the  same  thickness. 

Another  most  important  precaution  is  the  preservation  of  the 
proper  angle,  as,  in  order  to  establish  the  required  curve,  each  row 
must  incline  slightly  from  the  vertical — the  lower  end  of  the  bricks 
being  in  contact,  which  is  not  the  case  with  their  upper  extremities. 

The  establishment  and  preservation  of  the  proper  curvature  are 
facilitated  by  the  occasional  interpolation  of  a  longitudinal  row  of 
wedge  shaped  or  key-brick,  technically  called  "bullheads."  These 
are  usually  only  obtainable  made  from  fire-clay,  but  are  almcst 
indispensable  for  the  center  row  when  the  final  keying  of  the  arch 
is  effected.  Otherwise,  the  entire  row  of  key-brick  must  be  cut 
from  common  brick,  an  arduous  aud  imperfect  task. 

The  keying  is  a  matter  of  some  delicacy,  and  should  be  i:ieY- 
formed  by  a  single  workman,  who  should  select  or  cut  his  keys  of 
such  thickness  as  to  produce  a  uniform  moderate  pressure  throngh- 
ont  the  entire  distance,  no  more  force  being  exerted  to  drive  the 
key  into  place  than  can  be  easily  effected  by  a  light  mason's  ham- 
mer, using  an  intervening  block  of  wood  to  prevent  the  destruction 
of  the  brick. 

While  the  masons  are  thus  employed,  the  blacksmith  and  his 
helper  should  have  completed  the  buckstaves  and  tie-rods  from 
measurements  furnished  by  the  foreman  u  ason  as  the  work  pro- 
gresses, it  being  in  such  cases  easier  to  suit  the  length  of  the  tie- 
rods  to  the  completed  mason  work  than  to  pursue  the  opposite 

As  soon  as  the  arch  is  completed,  the  head  mason  aud  blacksmith 
should  proceed  to  the  ironing  of  the  furnace,  which,  with  the 
assistance  of  two  laborers,  should  be  completed  in  a  single  day. 

The  most  convenient  and  easily  obtained  buckstaves  in  many 
cases  are  old  iron  rails  of  full  size,  say,  80  pounds  to  the  yard. 
Properly  shaped  I-beams,  of  corresponding  strength,  are  about  15 
per  cent,  lighter.  The  tie-rods  may  consist  of  one  inch  round- 
iron  for  the  bottom  rod,  and  one  inch  and  a  quarter  iron  for  the 
npper  rods.  The  lower  rods  are  already  long  in  place,  and  through 
each  of  their  loops  should  now  be  slipped  one  of  the  upright  buck- 
staves,  cut  to  the  proper  length,  and  temporarily  wedged  into  the 
loop  to  keep  it  perpendicular. 


The  upper  tie-rods  may  be  made  the  same  as  the  lower,  with  a 
loop  at  each  end — the  necessary  tigliteniug  being  eUected  by  flat 
iron  wedges;  or  they  may  have  a  threaded  extremity  at  one  end 
passing  through  a  corresponding  hole  in  the  buckstaff,  and  fitted 
with  a  strong  nut;  or,  best  of  all,  a  small  ring  is  formed  at  one 
end  of  the  tie-rod,  through  which  slips  a  U-shaped  piece  of  round 
iron,  which  fits  against  the  buckstaff,  on  the  other  side  of  which  a 
piece  of  flat  iron,  pierced  with  two  holes  for  the  free  ends  of  the 
U  is  held,  these  ends  being  threaded;  a  nut  for  each  of  the  ends 
completes  the  apparatus,  and  presses  the  piece  of  flat  iron  tight 
against  the  upright.  This  is  a  simple  and  highly  satisfactory 
device,  and  avoids  the  disagreeable  process  of  wedging  in  the  oue 
ease,  or  of  punching  a  large  hole  through  a  narrow  rail  in  the 
other.  The  strain  is  distributed  over  two  bolts  and  nuts,  and  can 
be  instantaneously  increased  or  diminished;  nor  will  the  nuts  rust 
solid  into  place,  provided  they  are  saturated  with  oil  annually, 
and  slightly  turned,  to  free  them. 

Whatever  method  of  tightening  the  tie-rods  may  be  selected,  the 
process  of  ironing  or  anchoring  should  begin  with  the  first  tie-rod 
on  the  main  hoclji  of  the  furnace,  nearest  tlie  fire  end,  and  ])roceed 
systematically  toward  the  rear,  thence  returning  to  the  shorter 
transverse  rods  that  support  the  aroli  over  the  gi'ate,  and  termi- 
nating with  the  long  longitudinal  rods,  which,  for  convenience  of 
handling,  should  be  in  three  lengtiis,  connected  with  hooks  and 
eyes.  Up  to  this  time,  no  great  strain  should  be  put  upon  the 
rods,  everything  being  merely  brought  to  a  solid  bearing;  but  after 
all  are  in  place,  and  the  buckstaves  evened  both  vertically  and 
laterally,  the  rods  may  be  drawn  to  the  desired  tension,  the  skew- 
back  being  still  further  supported  by  a  bar  of  one  by  four-inch  flat 
iron,  or,  better,  an  iron  or  steel  rail,  let  in  flush  with  the  brick- 

This  is  largely  a  matter  of  experience,  and  being  of  vital  impor- 
tance should  receive  the  most  careful  attention  on  the  part  of  the 
builder,  as  too  lax  a  condition  of  the  rods  may  permit  the  entire 
falling  in  of  the  arch,  while  the  contrary  fault  may  cause  a  positive 
buckling  and  elevation  of  the  same,  accompanied  with  a  general 
cracking  and  distortion  of  the  lateral  walls.  The  latter  accident, 
in  a  moderate  degree,  is  much  more  likely  to  occur  than  the  for- 
mer, owing  to  the  natural  tendency  to  overdo  a  measure  essential 
to  safety,  and  yet  not  exactly  defined. 

The  lateral  rods  should   be  tightened  until  they   begin,  when 


Struck  near  the  center  with  a  hammer,  to  vibrate  rapidly,  and  to 
be  but  little  depressed  when  stepped  upon.  (It  is  almost  needless 
to  say  that  none  of  the  upper  rods  should  touch  the  arch.)  A 
simnltaneons  examination  of  the  brick-work  forming  the  upper 
portion  of  the  side  walls  should  also  be  made,  as  it  is  here  that  the 
elfect  of  the  curving  of  the  buckstaff  from  too  great  tension,  and 
consequent  pressure  agaiust  the  mason-work,  is  first  visible 

The  extreme  limit  of  tension  is  reached  when  the  first  signs  of 
tliis  appear,  as  notiiing  can  be  gained  by  bending  the  uprights,  and 
if  the  latter  are  sufficiently  strong  and  applied  in  the  numbers 
shown  in  the  illustration,  tlie  arch  may  be  considered  perfectly 
supported.  All  the  rods  should  be  tightened  to  about  the  same 
extent,  although  it  must  be  remembered  that  the  great  length  of 
the  longitudinal  rods  may  prove  deceptive  in  estimating  their  ten- 
sion, it  being  impossible  to  tighten  them  to  such  a  degree  as  the 
shorter  lateral  ones. 

A  single  additional  precaution  is  recommended,  though  seldom 
practised  by  builders.  This  consists  in  breaking  up  a  few  thin 
roofing  slates  into  fragments  a  couple  of  inches  in  length,  and 
driving  these  with  moderate  force  into  wliatever  crevices  may  still 
be  found  in  the  surface  of  the  arcli. 

Some  twenty  or  thirty  pails  of  liquid  mud  are  now  poured  over 
the  arch,  and  the  process  repeated  as  it  dries  until  every  crack  and 
crevice  is  filled,  and  the  roof  rendered  completely  solid  and  air- 

The  wooden  center  on  which  the  arch  was  built  should  now  be 
removed  by  first  knocking  away  the  little  posts  that  support  it, 
using  a  light  stick  of  timber  as  a  battering-ram,  and  proceeding 
from  one  side  door  to  the  next  until  every  stick  and  batten  are 
removed.  They  should  be  stored  for  future  use.  Any  indications 
of  settling  on  the  part  of  the  arch  must  be  immediately  counter- 
acted by  tightening  the  tie-roils;  but  when  the  precautions  enu- 
merated have  been  carefully  observed,  this  can  never  occur. 

The  length  of  time  the  completed  furnace  may  now  stand  un- 
touched with  advantage  to  the  mason-work  is  only  limited  by  the 
requirements  of  the  business,  which  almost  invariably  demand  its 
being  put  in  commission  at  the  earliest  possible  moment.  Under 
such  circumstances  a  smoldering  fire  of  large  logs,  knots,  or  any 
slow-bnrning  waste  material,  should  at  first  be  kindled  on  the  floor 
of  the  ash-pit,  the  grate-bars  not  being  put  in  place  until  the 
masonry  surrounding  the  fireplace  is  partially  dried. 


lu  twelve  or  eighteen  hours  tlie  lire  is  elevated  to  its  proper 
place,  and  with  a  nearly  closed  ash-pit  door  ajul  partially  lowered 
damper,  the  j)i"ocess  of  drying  proceeds  gently  and  without  that 
violent  generation  of  steam  and  vapor  that  is  sure  to  be  accom- 
panied by  extensive  fissiiring  of  the  brick-work  and  permanent 
weakening  of  the  entire  structure. 

A  most  careful  and  repeated  examination  of  the  condition  of  tie- 
rods  and  buckstaves  should  be  made  every  few  hours  from  the 
first  kindling  of  the  fire  until  the  furnace  has  attained  its  full 
heat,  and  may  be  supposed  to  have  expanded  to  its  utmost  limits, 
.although  it  may  be  a  month  or  more  before  all  evidences  of  move- 
ment cease.  The  first  indication  of  this  process  will  be  seen  in  the 
neighborhood  of  the  bridge  and  fireplace,  where  the  highest  tem- 
perature prevails.  A  bending  of  the  buckstaves,  combined  with  a 
pressing  in  of  the  skewback  line  and  an  increased  tension  of  the 
•cross-rods,  are  warnings  that  may  soon  be  followed  by  either  a 
complete  giving  way  of  some  portion  of  the  iron-work,  or  more 
frequently  by  a  bodily  upheaval  of  the  arch  and  general  Assuring 
of  the  brick-work  unless  relieved  by  diminishing  the  strain  to  a 
corresponding  degree.  This  process  of  loosening  must  be  extended 
to  the  entire  iron-work  of  the  furnace,  and  continued  as  long  as 
necessary,  the  tension  being  again  increased  if  the  furnace  is  ever 
allowed  to  cool  down  to  any  considerable  degree — an  operation 
more  destructive  to  it  than  many  months  of  ordinary  wear. 

While  the  apparatus  is  thus  gradually  being  brought  into  proper 
heat,  the  sheet-iron  hopper  should  be  suspended  from  timbers  rest- 
ing upon  the  trussed  beams  of  the  building.  It  should  be  strongly 
constructed  and  well  braced,  and  provided  with  a  stout  lever,  easily 
:accessible  to  the  operator  when  standing  upon  the  floor  of  the 
building.  A  track  running  transversely  to  the  row  of  calcining- 
furuaces,  and  parallel  with  the  longitudinal  axis  of  the  building, 
renders  these  hoppers  easily  accessible  to  the  car  in  which  each 
weighed  charge  of  ore  is  brought.  The  car  is  provided  with  a 
dumping  arrangement,  so  that  it  easily  and  completely  empties 
itself  into  the  furnace  hopper.  The  laborer  who  weighs  and  trans- 
ports the  charges  can  supply  six  furnaces,  provided  everything  is 
arranged  as  herein  described,  or  in  a  similarly  judicious  manner. 

The  outfit  of  tools  may  now  also  be  prepared,  and  should  consist, 
for  each  four-hearth  calciner,  of  0  rabbles,  4  inches  by  10  inches 
and  12  tol4feet  long;  6  paddles,  8  inches  by  12  inches  and  14  feet 
long;  4  door-hooks,  to  handle  the  sheet-iron  working-door;  1  long' 


hooked  and  pointed  iron  poker  for  wood,  or  an  ordinary  coal  poker. 
if  coal  is  used;  2  ordinary  long-liandled,  square-pointed  shovels;  1 
scoop-shovel  (for  coal). 

The  irou  rollers,  usually  employed  as  rests  for  the  long  tools  at 
each  working-door,  soon  lose  their  shape  and  cease  to  revolve.  It 
is  better,  therefore,  to  provide  merely  a  smooth  iron  bar,  which, 
if  kept  well  soaped,  renders  the  handling  of  the  tools  as  easy  as 
any  of  the  more  expensive  devices. 

When  available,  a  free-burning  semi-bituminous  coal  forms  the 
most  economical  fuel  for  calcining  purposes,  but  should  always  be 
burned  upon  a  comparatively  shallow  grate,  instead  of  using  the 
deep  clinker  bed,  so  suitable  to  the  smelting  process.  At  the  com- 
paratively low  temperature  suited  to  calcination,  the  generated  gas 
does  not  burn  perfectly,  and  a  great  waste  of  fuel  occurs.  Coal 
siiould  be  fed  at  short  intervals — from  30  to  45  minutes — in  quan- 
tities seldom  exceeding  50  pounds.  When  wood  is  cheap,  nothing 
can  excel  it  as  a  fuel  for  calcining  purposes,  its  long,  hot,  non- 
reducing  flame  being  peculiarly  suited  to  the  requirements  of  the 
process.  About  one  and  two-thirds  cords  of  hard,  or  two  cords  of 
soft  wood  are  commonly  considered  equal  to  2,240  pounds  of  fair 
bituminous  coal. 


The  most  important  feature  of  a  chimney  is  its  foundation;  but 
it  is  at  this  very  point  that  a  great  saving  over  ordinary  practice 
may  be  effected  without  lessening  the  stability  of  the  superstruc- 

A  mere  increase  in  depth  below  the  loose  soil  forming  the  surface 
of  the  ground  does  not  add  in  the  slightest  to  the  value  of  the 
foundation,  after  a  proper  material  for  the  same  has  once  been 
reached;  and  as  this  occurs  in  the  greater  number  of  cases  within 
three  or  four  feet  of  the  surface,  the  frequent  practice  of  additional 
excavation  for  the  apparent  purpose  of  merely  g::ining  depth  is 
money  thrown  awav. 

After  removing  the  loose  surface  soil,  and  penetrating  below  any 
danger  of  frost,  in  the  greater  number  of  cases  no  advantage  would 
be  gained  bv  excavating  to  a  depth  of  50  feet,  unless  solid  bed-rock 
were  reached. 

Any  kind  of  gravel,  hard-pan,  or  even  soft  loam  or  sa«d,  if 
homogeueous,  will  answer  the  purpose  perfectly,  it  being  under- 


Stood  that  reference  is  here  made  to  au  ordinary  calciuer  or  smelter 
stack  not  exceeding  80  feet  in  height. 

In  the  case  oi  a  yielding  sand  bottom,  and  especially  if  the  line 
of  division  between  two  strata  of  varying  quality  happens  to  cross 
the  excavation,  it  is  well  to  form  a  solid  floor  to  the  pit  by  putting 
in  a  double  layer  of  3-inch  plank,  nailed  crosswise.  But  in  all 
ordinary  cases  the  hole  should  be  simply  filled  with  broken  stone, 
about  the  size  of  ordinary  road  metal.  This  material,  when  well 
rammed  into  phice  and  thoroughly  grouted,  by  pouring  in  a  suffi- 
cient quantity  of  mortar  composed  of  one  part  each  of  lime  and 
cement,  and  three  of  sand,  makes  a  foundation  infinitely  superior 
to  one  formed  of  a  few  large  stones,  the  slightest  settling  of  any 
one  of  which  will  throw  the  chimney  out  of  perpendicular. 

The  excavation  should  beat  least  three  feet  larger  in  every  direc- 
tion than  the  base  of  the  chimney,  and  the  stone-work  of  the  latter, 
laid  in  lime  and  cement,  may  cease  some  three  feet  below  the 
surface,  at  which  point  the  brick-work  usually  begins. 

If  a  smel ting-furnace  is  in  operation  in  the  immediate  vicinity, 
nothing  can  be  more  satisfactory  or  economical  than  the  following 
plan,  pursued  by  the  author  on  several  occasions: 

An  excavation  being  made  of  the  usual  size,  the  molten  slag 
from  the  smelting-furnace  is  wheeled  to  the  spot  in  the  usual  mov- 
able shjg-pots,  and  poured  at  once  into  the  hole,  which,  when  filled 
to  the  proper  height  with  the  fused  rock,  and  leveled  by  means  of 
little  clay  dams  along  the  edges,  so  as  to  present  a  smooth  surface 
for  the  masons  to  begin  on,  will  contain  a  solid  block  of  lava, 
weighing  many  tons,  and  as  immovable  as  a  ledge  of  rock. 

In  constructing  a  stack  we  have  to  determine  the  size  of  flue 
desired,  and  intimately  connected  with  the  same  is  tl^e  degree  of 
batter,  or  taper,  which  shall  be  given  to  the  structure. 

The  object  of  this  batter  is  twofold:  1.  For  appearance.  '2. 
For  the  sake  of  strength.  The  first  reason  may  be  entirely  neg- 
lected in  metallurgical  architecture,  and  experience  has  shown 
that,  within  the  limit  of  height  mentioned,  a  batter  of  one-eighth 
of  an  inch  to  the  foot  is  ample.  Nor  need  the  taper  be  begun 
until  the  stack  rises  above  the  roof,  as  that  portion  of  the  structure 
within  the  building  is  amply  protected  from  the  force  of  the  wind. 
By  thus  decreasing  the  amount  of  taper,  we  greatly  increase  the 
capacity  of  the  stack,  as  experience  shows  that  a  contraction  nf 
the  flue  in  its  upper  portion  is  accompanied  with  a  corresponding, 
dimitiutiou  cl  draught,  while  a  positive  enlargement  of  the  same 


toward  the  top  has  a  most  beneScial  iuflueuce.  This  latter  point 
is  gained  by  lessening  the  thiciiness  of  the  chimney  walls  as  they 
grow  higher,  while  the  outside  taper  remains  constant. 

All  calculations  and  formula?  regarding  the  necessary  size  of  any 
flue  for  a  given  duty  have  been  found  so  greatly  modified  by  rir- 
cumstauces — such  as  variations  of  internal  and  external  tempera- 
ture; humidity  of  atmosphere  and  state  of  barometer;  change  of 
Avinds,  etc. — that  it  is  found  safest  to  rely  upon  experience  and 
analogy;  and  after  beginning  with  a  much  larger  flue  for  safety, 
the  author  has  6ually  found  a  stack  42  inches  square  inside,  at  its 
narrowest  part,  and  05  feet  high,  to  possess  ample  capacity  for  two 
large  calciuing-furnaces  such  as  just  described.  It  is  proper  tc 
^idd  that  a  much  smaller  stack  will  produce  the  draught  usually 
con&.dered  as  quite  sufficient  for  the  calcining  process;  but  long- 
continued  experiment  has  shown  such  extraordinarily  favorable 
results,  as  regards  both  capacity  and  perfection  of  roast,  to  arise 
fro'n  greatly  increasing  the  ordinary  calciner  draught,  that  a  sharp 
and  powerful  draught  appears  as  essential  to  a  calciner  as  to  a 

For  this  reason,  also,  no  more  than  two  furnaces  should  be  led  into 
a  common  low  stack,  it  being  almost  impossible  properly  to  equal- 
ize the  admission  of  air  to  each  calciner,  and  to  produce  that  sharp 
and  vigorous  draught  so  essential  to  ra})id  oxidation,  and  especially 
to  the  conveyance  of  the  sheet  of  flame  and  heated  gases  over  the 
whole  length  of  a  4-hearth  calcining-furuace.  The  interposition 
of  dust-chambers,  or  preferably  of  large  flues,  filled  with  parallel 
rows  of  sheet-iron,  according  to  the  method  found  so  eflicient  and 
economical  at  Ems,  is  of  course  necessary,  and  should  be  present 
in  any  case.*  Limited  experiments  conducted  by  the  author  fully 
satisfy  him  of  the  great  benefit  to  be  derived  from  the  adoption  of 
this  economical  and  efficient  metliod  of  condensation. 

The  size  of  chimney  mentioned — 42  inches — will  answer  for  all 
elevations  np  to  5,000  feet  above  sea-level.  For  each  1,000  feet 
additional  height,  these  figures  should  be  increased  one  inch. 

For  a  calciner  chimney  of  this  size  and  65  feet  in  height,  the 
Avails  at  the  base  shoald  be  IT  inches  thick,  the  length  of  two  red 
brick,  no  fire-brick  being  needed,  as  the  gases  are  snflficiently 
cooled  by  their  passage  through  the  long  furnace  and  flue.  This 
thit'kness  is  maintained  for  a  height  of  25  feet  from  the  ground, 

"See  description  of  Ems  metbod  of  coiult-nsation,  b_v  Professor  Egleston,  in 
Transactions  of  the  American  Institute  of  Mining  Engineers,  XI.,  379. 


which  briugs  it  somewhat  above  the  roof  of  the  building.  At  this 
point,  the  external  batter  of  one-eighth  of  an  inch  to  the  foot  is 
begun,  and  an  internal  set-off  of  4  inches  is  taken;  thus  decreasing 
tiie  thickness  of  the  walls  to  13  inches,  and  enlarging  the  flue  to 
50  inches. 

This  constant  taper  is  maintained  by  the  employment  of  an 
ordinary  beveled  plump-bob,  which  obviates  any  trouble  or  calcu- 
lation. This  condition  of  affairs  is  continued  for  another  25  feet, 
during  which  distance  the  flue  is  contracted  to  a  size  of  about  44 
inches,  when  another  internal  4-inch  set-off  is  taken,  incieasing 
the  same  to  52  inches,  while  the  walls  are  diminished  to  8  inches. 

This,  being  continued  for  15  feet,  gives  the  full  height  of  Grt 
feet,  the  flue  at  the  top  being  still  48  inches  square,  or  G  inches 
larger  than  at  the  base.  No  ornamental  fluish  at  the  lop  should 
ever  be  allowed,  the  stack  either  being  surmounted  by  a  light  cast- 
ing to  hold  the  brick  in  place,  or  left  without  this  protection,  the 
iron  braces  being  usually  sufficient  to  prevent  the  loosening  of  the 
upper  rows  of  brick-work.  An  ornumeutal  cap  is  simply  a  source 
of  annoyance  and  danger,  and  should  never  be  permitted  in  a  stack 
devoted  to  the  passage  of  sulphurous  vapors. 

A  chimney  of  this  size  is  best  built  from  the  outside,  a  scaffold 
being  erected  by  placing  eight  stout  poles  about  the  base  of  the 
proposed  structure,  nailing  crosspieces  at  the  proper  height  for 
the  plank  staging,  and  thoroughly  bracing  the  uprights  by  boards 
nailed  diagonally  from  one  to  the  other. 

The  uprights  may  be  lengthened  out  almost  indefinitely  by  care- 
ful splicing,  and  as  the  stack  grows  higher,  new  crosspieces  are 
spiked  every  five  feet,  and  men  and  material  thus  maintained  at 
the  desired  elevation.  A  rope  and  bucket,  with  a  single  wooden 
block  fastened  to  the  railing  of  the  staging,  and  manipulated  prin- 
cipally from  the  ground  level,  form  the  most  economical  means  of 
elevating  the  requisite  material,  while  a  single  laborer  above  is  able 
to  furnish  four  masons  with  brick  and  mortar,  most  of  the  work 
being  done  from  below.  It  is  best  to  employ  four  masons,  so  that 
one  can  work  on  each  wall  of  the  stack,  and  their  position  should 
be  changed  twice  daily,  in  order  to  equalize  any  differences  in  the 
amount  of  mortar  used,  etc. 

Like  all  other  mason-work  that  is  to  be  exposed  to  heated  sul- 
phurous gases,  the  interior  portion  of  the  stack  must  be  laid  in 
clay  mortar  (ordinary  sticky  mud);  while  the  remainder  of  the 
structure  should  be  laid  in  lime  mortar,  on  account  of  its  superior' 


tenacity.  To  prevent  the  penetration  of  the  vapors  into  the  porous 
brick,  the  interior  of  the  flue  should  be  thoroughly  plastered  with 
clay  throughout  its  entire  extent. 

While  the  durability  of  a  chimney  of  this  description  is  largely 
dependent  upon  its  being  ironed,  it  is  still  more  dependent 
upon  its  not  being  ironed  too  stiffly.  A  stack  with  corners 
thoroughly  inclosed  in  stiff  angle  iron,  tightly  held  together 
with  frequert  braces,  may  fissure  and  give  out  in  a  few  years,  while 
a  similarly  built  chimney  containing  a  few  light  irons,  merely  to 
hold  the  brick-work  in  place,  will  last  twenty  years  or  more. 

This  is  the  result  of  personal  experience,  confirmed  by  the  obser- 
vations of  most  other  constructing  engineers,  and  is  especially  the 
case  in  countries  where  high  winds  and  violent  fluctuations  of 
temperature  are  prevalent. 

Eight  uprights  of  f-inch  by  f-inch  iron,  each  upright  being 
placed  about  4  inches  from  each  corner  of  the  stack,  and  passing 
through  rectangular  openings  cur  in  one-half  by  2-inch  flat  iron, 
which  latter  pieces  are  Jaid  in  the  brick-work  from  30  to  36  inches 
apart,  are  amply  tiufficient  for  the  purpose.  The  holes  must  be 
so  punched  that  the  uprights  can  be  wedged  tightly  against  the 
brick-work,  which  is  thus  held  in  place  even  after  the  mortar  has 
long  succumbed  to  the  combined  influence  of  the  roast  gases  and 
the  elements.  As  a  striking  example  of  the  accuracy  of  the  above 
remarks,  the  reverberatory  smelter  stacks  of  the  Detroit  Smelting 
Company's  copper-refining  furnaces  at  Lake  Superior  may  be  men- 
tioned, where,  on  building  a  strongly  ironed  stack,  they  found  it 
fissure  and  become  unsound  in  a  very  short  time;  whereas  their 
ordinary  stacks,  anchored  only  by  means  of  occasional  straps  of  flat 
iron  built  iiito  the  chimney  walls  and  bent  over  at  each  end,  have 
stood  for  fifteen  years  or  more  without  showing  crack  or  imper- 

A  row  of  headers  should  be  introduced  at  about  every  eighth 
course,  and  the  lower  portion  of  the  stack  into  which  the  two  cal- 
ciner  flues  enter  on  opposite  sides  should  be  divided  by  a  4-inch 
partition  wall  into  two  equal  compartments.  This  wall,  extending 
some  five  feet  above  the  entrance  flues,  serves  to  bend  each  current 
in  an  upward  direction,  and  thus  prevent  the  whirl  and  disturbance 
of  draught  resulting  from  the  meeting  of  two  opposing  currents. 

The  following  interesting  observation  has  been  communicated 
by  Messrs.  Cooper  and  Patch,  superintendent  and  chemist  of  the 
Detroit  Refining  Works: 


In  iiios'^  reverberatory  fiiruaces,  the  flue  enters  the  stack  at  soint! 
■distance  above  its  base,  and  consequently  there  is  a  cavity  inclosed 
by  the  chimney  walls,  of  greater  or  less  depth  below  the  em- 
bouchure of  the  flue.  When  this  apparently  useless  cavity  has 
become  filled  up  from  the  falling  in  of  the  stack  lining,  drippings 
from  the  molten  brick,  or  other  causes,  the  draught  at  once  suli'ers 
^ind  the  capacity  of  the  furnace  is  greatly  diminished. 

Whether  this  phenomenon  arises  from  the  loss  of  the  elastic  air- 
cushion  that  is  normally  present,  or  whether  there  is  some  other 
reason,  the  fact  remains,  and  although  the  observations  have  been 
confined  mostly  to  smelting-furnaces,  it  is  probable  that  a  calcining- 
furnace  may  be  affected  in  a  similar  manner,  and  therefore  in  all 
cases  where  a  horizontal  or  inclined  flue  enters  a  stack,  it  should 
be  so  constructed  as  to  leave  an  open  space  of  from  4  to  G  feet 
below  it.  This  need  not  communicate  with  the  outside  air  in  any 
way,  except  for  the  purpose  of  cleaning  the  stack  or  entering  it  for 

It  is  well  to  provide  every  high  stack  with  a  good  lightning  rod, 
properly  fastened  and  insulated. 

The  building  that  covers  any  considerable  number  of  calcining- 
furnaces  is  necessarily  of  great  extent,  and  should,  if  possible,  be 
built  of  very  light  and,  at  the  same  time,  fireproof  materials. 

Scarcely  anything  tills  these  requirements  so  thoroughly  as  a 
medium  grade  of  corrugated  iron.  This,  if  well  fastened  down, 
and  painted  every  three  or  four  years,  will  be  found  the  most 
economical  and  satisfactory  material  for  both  sides  and  roof  that 
is  yet  known.  If  the  number  of  furnac  s  under  a  single  roof  ex- 
ceeds two,  they  should  be  placed  at  right  angles  to  the  greatest 
length  of  the  building,  a  space  of  only  three  feet  being  left  between 
the  rear  end  of  the  furnace  and  the  corresponding  side  of  the 
building,  while  between  the  fire-box  and  the  lower  side  of  the 
building  there  should  be  ample  room  for  a  driveway  for  the  con- 
veyance of  fuel,  as  well  as  for  a  railroad  parallel  to  the  same  and 
close  to  the  wall,  over  which  the  calcined  ore  may  conveniently  be 
dumped  into  a  paved  and  roofed  inclosure  on  a  level  as  low  as  the 
circumstances  of  the  case  permit.  The  16-foot  calciners  should 
be  separated  by  spaces  of  at  least  fourteen  feet. 

As  the  main  building  for  these  long  calcining-furnaces  must  be 
from  80  to  V)0  feet  in  width,  it  is  often  the  practice  to  support  the 
cross-beams  on  posts  that,  if  properly  placed  close  to  the  furnace 
and   midway  between    the   working   openings,   need   not  interfere 


with  tlie  long  tools  in  use.  But  there  is  no  difficulty  in  construct- 
ing trusses  to  support  a  roof  of  this  size  without  the  aid  of  posts^ 
nor  need  the  expense  he  much  greater.  The  principal  difficulty  is 
encountered  in  raising  these  immensely  long  and  heavy  "bents;" 
but  this  inay  be  entirely  obviated  by  constructing  a  series  of  cheap 
scaffoldings,  and  putting  them  together  piece  by  piece,  instead  of 
attempting  to  raise  the  entire  *' bent"  bodily.  The  ridge  of  the 
roof  should  be  surmounted  by  a  continuous  ventilator  throughout 
its  entire  extent.  The  details  of  this  work  may  be  intrusted  to 
any  experienced  carpenter, 


The  following  estimates  of  cost  are  taken  from  notes  that  cover 
the  construction  of  a  considerable  number  of  large  calciuing- 
furnaoes,  and  being  given  without  alteration  or  omissions,  except- 
ing the  necessary  reduction  to  our  assumed  standard  of  costs, 
should  furnish  reliable  figures  on  which  to  base  future  plans: 


Excavation— 45  days  at  $1.50 $67.50 

Removal  of  material  excavated 35.00 

Superintendence  and  miscellaneous 24.00       $126.50 

Foundation  walls— 1,840  cubic  feet. 

2,000  slag-brick  at  2  cents 40.00 

20  days  stone-mason  and  helpers 120.00 

Materials  for  mortar 28.00 

Labor  on  same  and  utensils 16.00 

Miscellaneous   labor 12.00 

Superintendence 15.00       $231.00 

Brick-work  on  furnace  proper. 

2,420  cubic  feet,  say  .50,000  red  bvick  at  $8 400.00 

7,500  fire-brick  at  $40 300.00 

Lime  and   sand 137.00 

4  tons  fire  clay  at  $8 32.00 

8  tons  brick-clay  at  $1.50 12.00 

32  loads  sand  at  $1.50 48.00 

112  days"  brick-masons'  labor  at  $4 448.00 

112  days'  ordinary  labor  at  $1.50 168.00 

3  days'  carpenters'  labor  at  $3 9.00 

Miscellaneous  labor 35.00 

8  days,  blacksmith  and  helper 40.00 

Materials  consumed  by  same 8.00 

i^upprintendence 112.00    $1,749.00 


Carried  forward ,.,.  $1,749.00 

Iron  Avork. 
66  buckstaves  (old  rails),  6|  feet  long,  80  pounds,  at  If  cents 

per  pound 85.80 

Tie-rods  and  loops,  2,056  feet,  li-inch  round  iron  =  8,327 

pounds,  at  2  cents 166.54 

Flat-iron   for  skewback,  grates,  etc.  =  2,064  pounds,  at  2 

cents 41.28 

16   cast   frames  and   doors,    at   156   pounds   each  =2,496 

pounds,  at  2i  cents .       62.40 

Fire-doors  and  other  small  castings 16.50       $372.52 

Nuts  and  bolts 6.25 

Short  flue  with  damper,  and  one-half  cost  of  stack 364.00 

Grading  and  miscellaneous 47-50 

Tracks  for  feed  and  discharge  of  ore 62.40 

Set  tools,  complete,  as  per  former  schedule,  1,250  pounds, 

at  2  cents 25.00 

Labor  on  same 18.00 

One  iron-ore  car  (list  price) 85.00 

Grand  total ,  , $3,087.17 

The  repairs  oa  a  thoroughly  built  calciner  shonld  be  nothing  for 
the  first  three  years;  for  the  succeeding  seven  years  they  will 
average  3  per  cent,  per  annum  on  its  first  cost,  while  from  its 
tenth  to  its  fifteenth  year,  5  per  cent,  per  annum  will  probably  be 
expended  in  renewing  the  hearth  and  roof  once  and  patching  the 
furnace  in  various  places. 

After  fifteen  years  of  constant  usage,  it  is  cheaper  to  build  a 
new  furnace  than  to  keep  the  old  one  in  repair;  but  few  metallur- 
gical enterprises  in  this  country  require  to  provide  for  a  period 
longer  than  the  above. 


In  roasting  a  heavy  pyritous  ore  for  subsequent  reverberatory 
smelting,  as  at  Butte,  Montana,  where  concentrates  with  40  per 
cent,  sulphur  require  to  be  roasted  down  to  7  or  8  per  cent,  sul- 
phur, a  large  calciner,  properly  and  energetically  managed,  will 
put  through  13  tons  of  ore  per  24  hours  with  a  consumption  of  2 

*  As  most  of  the  costs  of  calcining  in  this  work  are  based  on  Montana  or  Col- 
orado prices,  the  estimate  of  th&cost  of  calcining,  as  given  in  earlier  editions, 
has  been  changed  to  correspond  with  the  other  calcining  estimates  It  lias  also 
become  customary  to  burn  more  coal  in  calciners  than  formerly,  and  drive 
them  harder,  thus  increasing  the  capacity  of  the  furnace  and  often  the  cost 
per  ton  of  material  roasted. 


tous  of  slack  coal  and  with  the  services  of  four  raeu  workiug  12- 
hoiir  shifts  in  two  gangs. 

The  following  estimate  shows  the  minimum  expenses: 


Two  tons  slack  coal  at  |3.50 $7.00 

Four  furnace-inen  at  $3.50 14.00 

One-fourtb  weiffhiuau  at  $3.00 75 

Repairs,  lights,  and  miscellaneous 60 

Proportion  of  foreman 50 

Interest  on  $4,000  at  6  per  cent,  per  annum 66 

Total $23.51 

Or  about  $1.81  per  ton  of  raw  ore. 


The  variety  of  reverberatory  calciner  known  as  the  muffle  furnace 
is  now  seldom  used  by  the  copper  smelter,  as,  except  for  purposes 
of  acid  manufacture,  it  possesses  few  advantages  above  the  ordinary 
hearth  variety,  and  in  case  this  branch  of  metallurgy  is  also  prac- 
tised, some  of  the  newer  forms  of  automatic  furnaces  have  dis- 
placed the  muffle.  The  high  cost  of  construction  and  greater 
consumption  of  fuel  are  also  adverse  to  its  employment,  and 
although,  from  its  gentle  and  regular  heat,  it  possesses  decided 
advantages  in  the  treatment  of  easily  fusible  substances,  it  is  rather 
suited  to  the  calcination  of  matter  containing  much  lead,  or  of 
pyrites  with  salt,  as  in  the  Henderson  process,  none  of  which 
operations  come  within  the  scope  of  this  treatise. 

An  easily  fusible  ore  can  be  very  efficiently  protected  from  the 
tierce  heat  of  the  first  hearth  of  an  ordinary  calciner  by  the  con- 
struction of  a  4-inch  curtain  arch,  covering  one-third  or  more  of 
its  surface  from  the  fire-bridge  onward,  though  such  a  precaution 
is  seldom  necessary,  excepting  in  the  case  of  matte  calcination, 
which  requires  but  slight  modifications  of  the  roasting  process  as 
applied  to  ordinary  sulphide  ores. 


These  also  are  extensively  and  advantageously  used  for  the 
chloridizing  of  silver  ores,  having  a  considerable  capacity,  and 
effecting  a  thorough  chloridization  at  a  very  moderate  cost.  They 
consist  essentially  of  a  horizontal  or  inclined  brick-lined  iron  cylin- 
der, revolved  slowly  by  gearing,  and  having  a  fireplace  at  one  end 
— or  at  both  ends,  used  alternately. 



The  cyliuder  with  continuous  discbarge  that  has  been  most 
largely  used  for  the  oxidizing  roasting  of  copper  ores  or  products 
is  the  White- Hoioell  and  its  imitations. 

This  is  a  slightly  inclined  cyliuder  of  small  diameter  in  propor- 
tion to  its  length,  is  lined  with  brick,  and  is  cradled  between  sup- 
porting rollers,  being  slowly  revolved  by  means  of  pinion  and 
spur-gear.  Four  longitudinal  ridges  of  brick-work  project  slightly 
from  the  inner  lining  at  intervals  of  90  degrees,  and  lift  the  ore 
until  it  falls  back  through  the  flame  in  a  thin  stream,  and  is  con- 
tinuously discharged  ac  the  tire-box  end  of  tne  cylinder. 

The  amount  of  ore  treated  is  determined  by  the  speed  at  which 
the  furnace  is  revolved,  and  the  angle  of  inclination  at  which  it 
is  set. 

It  is  largely  used  for  the  chloridizing  roasting  of  silver  ores,  but, 
in  spite  of  its  many  seeming  good  points,  has  never  been  very 
popular  among  copper  men. 

At  the  works  of  The  Cape  Copper  Company,  Limited,  at  Briton 
Ferry,  and  The  Messrs.  Elliott's  Company's  Works  at  Pembrey, 
South  Wales,  I  saw  several  cylinders  of  the  above  general  pattern 
running  on  76  per  cent,  white  metal.  The  cylinders  were  7  feet 
in  diameter  and  60  feet  long,  having  an  inclination  of  5|-  inclies. 
They  made  8  revolutions  per  hour,  and  calcined  about  22,000 
pounds  of  the  metal  per  24  hours  down  to  1  per  cent.  to3  per  cent, 
sulphur.  The  consumption  of  coal  was  2,000  pounds  in  the  fire- 
box, besides  the  power. 

The  matte,  crushed  through  a  screen  with  three  meshes  to  the 
linear  inch,  was  conveyed  into  an  iron  hopper  at  the  cold  end  of 
the  furnace,  whence  it  flowed  by  gravity  into  the  cylinder  through 
a  two  and  a  quarter  by  half-inch  slot  in  the  floor  of  the  hopper. 
The  cylinder  required  one-fourth  of  a  laborer's  time  to  fire  and 
remove  ashes,  and  a  small  lad  to  watch  the  feeding,  oil  the 
machinery,  etc. 

The  cost  per  2,000-pound  ton  of  matte,  taken  from  the  results 
of  several  years'  running,  and  allowing  for  repairs  and  interest  on 
investment,  was  reported  to  me  as  33  cents.  This  did  not  include 
the  handling  and  re-roasting  of  the  flue-dust.  These  costs,  natu 
rally,  were  based  on  Swansea  prices  for  coal  and  labor,  but  they 
are  extremely  low,  and  are  interesting  as  showing  the  ease  with 
which  certain  sorts  of  matte  can  be  calcined  in  continuous  cylinders. 


The  continuous  discharge^  muffle  cylinder-calciner  of  James 
Douglas  contains  a  heavy,  central  tile-flue,  supported  by  four  slotted 
tile  partitions.  The  products  of  combustion  are  thus  carried 
direct  into  the  chimney,  without  ever  mingling  with  the  roast 

Mr.  Douglas  invented  this  apparatus  for  the  calcination  of 
jiyrites  fines,  in  order  to  obtain  pure  and  concentrated  sulphurous 
acid  fumes  for  the  Hunt  and  Douglas  process  of  wet  copper  extrac- 
tion, but  it  has  proved  a  rapid  and  efficient  calcining  furnace  as 

After  the  heavy  interior  mass  of  brick-work  has  become  once 
thoroughly  heated,  the  cylinder  will  do  good  work  without  the 
aid  of  carbonaceous  fuel.  Its  capacity  is  largely  influenced  by  the 
amount  of  air  admitted  to  the  roasting  chamber;  and,  as  the  pri- 
mary mission  of  this  apparatus  is  to  furnish  a  supply  of  concen- 
trated sulphurous  acid  gas,  its  actual  roasting  capacity  has  always 
been  seriously  handicapped.  Its  capacity  is  6  to  1"2  tons  of  pyrites 
fines  per  24  hours,  roasted  down  to  about  3  per  cent,  sulphur. 


At  the  present  time,  the  modified  and  improved  BrilcJcner^s 
cylinder  stands  pre-eminent  as  the  most  satisfactory  and  econom- 
ical of  all  revolving  cylinders  for  pulverized  ore.* 

The  large  cylinders,  as  now  made,  are  8  feet  6  inches  in  diameter 
by  18  feet  6  inches  long. 

They  are  lined  witii  one  thickness  of  good  red  brick,  though 
doubtless,  where  fire-brick  are  cheap,  it  will  pay  to  use  them,  as 
they  withstand  the  mechanical  wear  and  tear  much  longer  than 
red  brick;  owing  to  the  care  bestowed  upon  their  manufacture, 
they  are  much  more  regular  in  shape,  thus  forming  a  tighter  and 
more  perfect  circle  inside  the  iron  shell,  the  strength  of  which  can 
be  still  more  increased  by  having  the  brick  molded  to  order  to  snit 
the  inner  circle  of  the  cylinder.  Again,  they  are  much  more  dura- 
ble when    exposed   to   dampness  than   are  red    brick,   which   are 

♦In  describing  this  furnace,  it  would  be  unjust  not  to  mention  Messrs. 
Fraser  &  Chalmers,  of  Chicago,  whose  energv  in  introducing  it,  and  in  going 
to  great  trouble  to  modify  and  improve  it.  and  adapt  it  to  the  desulphurization 
of  copper  ores  and  concentrates,  has  earned  the  gratitude  of  a'.l  oopper-smelters. 
To  Mr.  W.  R.  Eckart.  the  profession  is  indebted  for  the  detailed  drawings  of 
the  cylinders  now  giving  such  satisfaction  at  the  Anaconda  Works,  in  Montana. 


quickly  destroyed  if  dripping  concentrates  are  fed  into  the  red-hot 

But  even  where  red  brick  are  used,  the  lining  lasts  about  18 
mouths  when  properly  pat  in,  and  as  this  is  the  principal  cost  of 
repairs  during  the  first  few  years,  it  is  evident  that  it  innst  be 
very  small. 

As  will  be  seen  in  the  accompanying  perspective  sketch  of  the 
furnace,  it  has  a  double-snouted  feed-hopper,  with  two  feeding- 
holes,  and  two  others  opposite  them,  halfway  around  the  circum- 
ference of  the  shell,  so  that  it  can  be  discharged  without  much 
loss  of  time.  It  is  best,  of  course,  where  possible,  to  discharge 
the  roasted  ore  directly  into  the  reverberatory  smelting  furnaces, 
if  such  are  used,  or  into  an  adjacent  vault,  where  the  heat  will  not 
be  rapidly  dissipated.  But  in  works  where  the  calcined  ore  must 
be  first  cooled  down  before  going  to  the  smelter,  the  cooling  ar- 
rangements must  be  of  large  capacity  to  handle  the  heavy  charge 
of  ore  employed. 

In  the  improved  cylinders,  the  fire-box  is  really  a  car,  running 
on  a  track  at  right  angles  to  the  longitudinal  direction  of  the 
cylinders,  and  having  a  short  flue  in  one  side  that  comes  exactly 
opposite  the  throat  of  the  furnace.  In  this  way,  the  fire-box  can 
be  run  opposite  a  cylinder,  which  contains  a  fresh  charge,  and  fired 
on  until  the  sulphur  is  fairly  kindled.  Then  the  movable  fire-box 
may  be  wheeled  along  to  a  neighboring  cylinder,  and  the  first  one 
left  to  complete  the  combustion  of  the  sulphur  with  a  free  access 
of  air,  and  undisturbed  by  the  reducing  gases  that  pass  through 
an  ordinary  grate.  After  the  combustion  of  the  sulphur,  it  is 
necessary  for  a  perfect  roast  to  again  connect  the  fire-box  with  the 
cylinder  and  supply  a  little  extraneous  heat  to  complete  the  de- 
composition of  the  sulphates. 

It  is  estimated  that  two  horse-power  are  required  to  drive  a 
charged  cylinder  at  average  speed.  The  size  and  weight  of  the 
ore-charge  varies  greatly  with  its  quality,  percentage  of  sulphur, 
specific  gravity,  etc. 

These  results  were  obtained  at  the  works  of  The  Anaconda 
Mining  Company  of  Montana,  where  156  of  the  cylinders,  8  feet 
by  18  feet,  are  in  operation.  The  following  results  comprise  the 
work  of  four  weeks  (28  days).  The  ore  calcined  consists  mainly 
of  coucentrates  containing  36.4  per  cent,  sulphur  and  16  per  cent, 
silica  and  is  roasted  down  to  8  per  cent,  sulphur. 

In  28  davs  a  cvlinder  treats  341  tons  of  drv  ore  or  12.186  ton? 

0  UI 








per  day.  It  uses  2.95  cords  of  wood  (378  cubic  feet)  per  day, 
costing  $10.26,  or  84|^  cents  j)er  ton  of  ore.  This  wood  conld  be 
unprofitably  replaced  by  2.63  tons  inferior  coal,  with  20  per  cent, 
ash;  or  by  1.625  tons  of  better  coal,  with  10  per  cent.  ash.  One 
laborer,  at  $3  per  shift  of  12  hours,  attended  three  furnaces,  cost- 
ing per  day  per  furnace,  12,  or  16.4  cents  per  ton  of  ore. 

his  makes  11.01  per  day,  to  which  must  be  added  the  expense 
of  power,  repairs  (small),  recrushing  and  recalcining  lumps  and 
rehandling  flue-dust  (large),  oil,  lights,  foreman,  and  interest  on 
plant,  $6,000.  This  brings  the  total  cost  to  about  ll.-iO  per  ton 
of  ore. 

By  the  courtesy  of  the  Chicago  Iron  Works,  I  am  enabled  to 
present  some  excellent  drawings,  showing  the  details  of  the  Bruck- 
ner cylinders  manufactured  by  them,  and  which  are  doing  very 
satisfactory  work  in  Montana  and  elsewhere.     (See  Plate  I.). 

Direct  statements  from  those  who  are  using  them  show  that  my 
estimate  of  a  saving  of  30  to  40  per  cent,  of  the  costs  of 
calcining  by  using  these  large  cylinders  in  lieu  of  hand-calcining 
furnaces  is  by  no  means  excessive,  and  in  some  instances  does  not 
represent  the  full  amount  saved. 



Automatic  hearth -furnaces  seem  to  oflEer  peculiar  advantages  as 
regards  capacity  in  proportion  to  first  cost,  and  ease  of  mauage- 
nieut.  They  are  also  used  for  the  roasting  of  leady  mattes  and 
other  material  that  is  inclined  to  sinter.  They  seem  peculiarly 
suited  to  roasting  pyritic  gold  ores  and  concentrates,  previous  to 
their  treatment  by  chlorination;  nor  can  I  see  why  they  could  not 
be  changed  into  muffle-furnaces,  that,  considering  the  space,  labor, 
and  plant  saved,  would  roast  pyrites  for  sulphuric  acid  manufac- 
ture more  economically  than  any  of  the  burners  at  present  in  use. 

The  most  prominent  furnaces  of  this  description  now  before  the 
public  are: 

The  O'Harra  furnace  with  certain  modifications  by  Allen  and 
by  Brown. 

The  Pearce  turret  furnace. 

The  improved  Spence  (Keller-Gaylord-Cole)  furnace. 

The  Brown  horseshoe  furnace. 

The  Spence  autonatic  desulphurizer. 

To  these  may  be  added  two  furnaces  that  are  solely  engaged  in 
roasting  zinc-blende  ores,  but  that  possess  features  of  interest  to 
the  copper  metallurgist,  viz. : 

The  Matthiessen  &  Hegeler  Zinc  Company's  furnace,  working 
at  La  Salle,  Illinois. 

Blake's  revolving  hearth  calciner,  at  Shullsburg,  Wisconsin. 

The  0' Harra  furnace  consisted  originally  of  two  long  hearths, 
one  above  the  other,  through  which  plows  were  continually  drag- 
ged by  means  of  a  chain  which  obtained  its  motion  from  grooved 
pulleys  over  which  it  ran.  The  chain  and  plows  were,  and  still 
are,  cooled  by  running  for  some  distance  outside  of  the  furnace  on 
their  way  from  the  lower  to  the  upper  hearth.  The  hearths  con- 
tained a  continuous,  longitudinal  groove,  not  for  the  chain  to  run 
in,  as  is  stated  by  Schnabel  and  certain  other  writers,  but  for  the 
protection  of  the  chain  in  case  of  shut-downs,  whose  probable  fre- 

lIB^d    S  MHOr 


qaeucy  and  extent  were  evidently  very  apparent  to  the  inventor. 
During  such  delays  the  tension  was  relaxed,  and  the  chain  was 
supposed  to  subside  into  the  groove.  The  hearths  were  heated  by 
a  sufficient  number  of  external  fireplaces  along  their  sides. 

While  the  capacity  of  this  furnace  was  large,  and  the  roasting 
satisfactory,  the  repairs  and  delays  were  excessive.  The  chain  and 
plows  riding  on  the  hearth,  constantly  gave  way  and  wore  out. 
The  plows  tore  up  the  hearth  and  dragged  it  to  the  front,  and  the 
life  of  a  furnace  scarcely  reached  eighteen  months. 

Allen  effected  a  radical  improvement  by  laying  iron  tracks 
through  the  hearths,  and  mounting  the  chain  and  plows  on 
wheeled  carriages. 

Brown's  modification  consisted  in  partitioning  off  little  corridors 
on  either  side  of  the  hearths,  in  which  the  tracks  were  laid,  and 
through  which  carriages  ran,  supporting  the  chain,  and  especially 
the  arm  to  which  the  plows  were  attached,  one  end  of  this  arm  be- 
ing fastened  to  the  carriage,  while  the  other  extremity  projected 
through  the  partition  wall  of  the  corridor  into  the  furnace.  It  is 
evident,  therefore,  that  Brown  had  to  use  two  chains,  and  two  sets 
of  rabbles,  their  arms  nearly  meeting  at  the  center  line  of  the 
hearth.  It  is  also  evident  that  there  had  to  be  a  continuous  slot 
in  the  partition  wall,  to  permit  the  travel  of  the  rabble-arm. 

This  modification  has  not  been  so  entirely  successful  as  its  in- 
genuity would  seem  to  deserve.  The  main  difficulty  has  been  the 
tendency  of  the  partition-tiles  or  castings  to  sag  or  loosen,  and 
obstruct  the  continuous  slot,  through  which  the  rabble-arm  pro- 
jects into  the  hearth.  This  trouble  has  been  remedied  by  Brown, 
and  independently  by  the  Argo  metallurgists,  but  a  mere  partial 
partitioning  oflE  of  the  hearth  does  not  seem  to  be  a  sufficiently 
perfect  means  of  protecting  the  tracks,  carriage  and  chain  from 
the  heat.  Those  who  are  running  O'Harra  furnaces  claim  that 
the  chain,  track,  etc.,  might  about  as  well  be  in  the  hearth  proper, 
as  it  was  before  Brown's  modification,  and  most  of  the  O'Harra 
furnaces  are  run  in  this  manner.  Brown  is  entitled  to  great 
credit,  however,  for  showing  us  the  use  of  a  continuous  slot  trav- 
ersed by  a  rigid  rabble-arm. 

The  Allen-O'Harracalciners  at  Butte  have  two  hearths,  each  9  by 
90  feet,  traversed  by  six  plows,  making  a  complete  circuit  in  3|  min- 
utes. They  roast  highly  pyritic  concentrates  containing  about  19 
per  cent,  copper,  18  per  cent,  silica,  and  40  per  cent,  sulphur,  down 
to  8  per  cent,  or  9  per  cent,  sulphur.     A  considerable  proportion 


of  the  concentrates  are  coarse,  anJall  are  wet.  They  lose  26  pel 
cent,  of  their  dry  weight  hy  calcination. 

The  weakest  point  of  even  the  improved  O'Harra  furnace  is  its 
heavv  repair  bill.  This  is,  to  a  considerable  extent,  unavoidable 
in  a  furnace  where  the  track,  carriages,  and  chain  are  all  exposed 
to  the  flame  and  to  the  red-hot  sulphides,  and  where  their  exist- 
ence is  entirely  dependent  upon  the  judgment  and  care  of  the  fire- 
men. But  both  Allen  and  Bellinger  of  Butte,  and  Fraser  & 
Chalmers  of  Chicago,  have  introduced  modifications  that  consider- 
ably lessen  the  cost  of  repairs.  The  wear  on  the  carriage-wheel 
bearings  is  rendered  unimportant  by  the  employment  of  cheap, 
renewable  bushings.  The  chain  has  always  been  one  of  the  most 
costly  portions  of  the  furnace,  for  though  made  of  hand-welded 
steel  links  it  is  apt  to  give  way  by  opening  at  the  welds.  Chains 
have  latelv  been  made  consisting  of  solid  steel  drop-forgings  for 
the  alternate  links,  these  being  connected  by  steel  Ds,  one  long 
tongue  of  which  is  put  through  an  eye  and  bent  over,  so  that  there 
is  no  weld  in  the  entire  chain.  The  consumption  of  fuel  has  been 
considerably  reduced  with  great  benefit  to  the  furnace  and  machin- 
ery, and  without  prejudice  to  the  roast.  The  cost  of  erection  has 
also  been  greatly  diminished,  while  the  furnace  is  stronger  and 
more  durable. 

The  improved  Allen-O'Harra  calciner  is  shown  in  Plate  II. 
The  ore  is  fed  automatically  from  the  hopper  A  on  to  the  upper 
hearth  B,  and  is  gradually  moved  by  the  plows  toward  the  further 
end  of  the  hearth,  where  it  drops  through  the  slot  C  on  to  the 
lower  hearth  T).  It  thence  traverses  the  lower  hearth  until  it 
reaches  the  discharge  E.  The  chain  is  driven  by  the  sprocket- 
wheel  F,  on  the  shaft  G,  and  is  kept  taut  by  the  wheel  H  in  the 
sliding  frame  I,  which  is  provided  with  a  weight,  J.  Six  sets  of 
plows,  K,  are  attached  at  equal  intervals  to  the  chain.  They  are 
carried  on  wheels  running  on  the  track  L.  The  chain  is  also  sup- 
ported by  simple  trucks  M,  midway  between  the  plow-carriages. 
It  will  be  noticed  that  the  vanes  on  the  separate  halves  of  the  same 
plow  turn  furrows  in  opposite  directions;  also,  that  the  same  plow 
on  the  upper  floor  turns  furrows  in  a  direction  opposite  to  its  fur- 
rows on  the  lower  floor,  and  that  each  plow  turns  furrows  in  a 
direction  contrary  to  those  made  by  the  plow  preceding  it.  A 
vane  set  to  turn  a  furrow  toward  a  guide-rail,  or  wheel,  is  fastened 
to  the  plow-shaft  at  some  distance  to  the  rail  and  wheel,  so  as  not 
to  cover  the  rail,  nor  to  throw  ore  into  the  path  of  the  wheel.     A 

;  ;   i  N 



vaue  set  to  turn  a  furrow  away  from  a  guide-rail,  or  wheel,  is  fas- 
tened on  the  plow-shaft  close  to  the  rail  and  wheel,  so  as  to  turn 
the  ore  away  from  them.  The  arrangement  of  the  vanes  on  the 
separate  halves  of  the  same  plow,  by  which  they  turn  furrows  in  an 
opposite  direction,  balances  the  tendency  of  the  plow  to  be  forced 
off  the  track  on  the  side  opposite  to  the  direction  of  the  furrows, 
which  it  would  have  if  the  furrows  were  all  turned  in  the  same 
direction.  The  hearths  are  closed  ai  each  end  by  horizontal  turn- 
stile doors  N,  actuated  by  the  moving  carriages.  The  cooling 
space  0  for  chain  and  plows  is  23  feet  in  length.  The  grid  P  at 
the  driving-end  of  the  furnace  is  intended  for  convenience  in  re- 
pairing chain  and  plows.  There  are  five  pairs  of  fire-boxes,  three 
for  the  lower  hearth  and  two  for  the  upper,  though  only  one  or 
two  pairs  are  commonly  used.  The  doors  E  are  provided  with 
dampers  to  admit  air  to  the  hearth.  The  tie-rods  that  pass  through 
the  upper  and  lower  floors  are  protected  by  2-inch  pipes,  and  may 
thus  be  easily  renewed  if  burned  out. 

The  costs  of  calcining  in  this  furnace  can  be  best  studied  at  the 
Alleu-O'Harras,  at  Butte,  Montana,  as  it  is  here  that  they  are 
working  on  copper  ores  on  the  largest  scale,  and  it  is  here  that 
they  were  first  adapted  to  the  purpose. 

It  is  difficult  to  oifer  an  estimate  of  costs  that  shall  seem  fair  to 
both  the  partisans  and  the  detractors  of  this  furnace.  The  prin- 
cipal cause  of  this  difficulty  is  the  fact  that  the  most  important 
items  of  cost  may  be  made  to  vary  from  50  per  cent,  to  150  per 
cent.,  according  to  the  care  and  skill  exercised  by  those  in  charge 
of  the  furnace.  These  items  are  the  fuel  and  the  repairs.  It  is 
very  easy  to  fire  in  all  sets  of  fireplaces  and  burn  10  cords  of  wood 
per  day;  but  equally  good  results  are  now  obtained  by  firing  in 
only  one  set,  and  burning  but  3.2  cords  of  wood  jier  day. 

Again,  a  very  little  carelessness  in  regulating  the  heat  may 
damage  the  chain  and  running  gear  to  the  extent  of  $100,  or 
more,  in  a  very  short  time,  and  augment  the  repairs  to  an  excessive 
sum.  But  careful  firemen  can  be  found,  and  a  month's  observa- 
tion of  ten  o|  these  furnaces  convinces  me  that  there  is  no  occasion 
for  damaging  irregularities  or  serious  delays. 

I  think  the  following  table  of  costs  will  be  found  about  correct 
for  the  Allen-O'Harra  furnace,  when  run  with  the  regularity  and 
skilled  supervision  that  it  receives  at  the  works  of  The  Montana 
Ore  Purchasing  Company,  or  the  Butte  &  Boston  Mining  Company. 

The  9  by  90  feet  double-hearth  calciners  at  these  works  average 

204  M013EKN    COrPER   SMELTING. 

50  tons  each  of  couceutrates  per  24  hours.  Much  cf  this  mate- 
rial is  very  coarse,  some  8  per  ceut.  of  it  coming  from  the  roughing- 
jigs,  aud  barely  passing  a  2-iiich  ring.  The  Butte  pyrites  decrepi- 
tates to  a  certain  extent.  The  average  of  150  partial  analyses  of 
certain  of  these  concentrates  is: 

Copper 12.3  per  cent. 

Iron 31.9 

Sulphur 41 .2 

Silica 10.6 

Silver 0.012    "  (4.4  oz.  per  ton.) 

95.912     " 

The  following  table  shows  the  cost  of  roasting  these  concentrates 
down  to  8  per  cent,  sulphur,  at  the  rate  of  50  tons  per  day  (100,000 
pounds)  per  furnace.  In  these  works,  there  is  one  foreman  and 
one  weighman  per  shift  to  eight  furnaces.  One  tireman  per  shift 
attends  two  furnaces.  One  gallon  of  black  oil  at  14  cents  is  used 
per  24  hours  for  the  machinery  of  the  eight  furnaces. 

No  transportation  of  ore  to  or  from  calciners  is  included. 


50  TONS. 

Total  Cost 

Expense.  per  Ton. 

Labor — i  foreman  at  $4.00 

1  fireman  at 4.00 

J  weighman  at 3.00 

$5.75  11.5  cents. 

Fuel— 3.2  cords  wood  (410  cubic  feet)  at  $4.70  per  cord.  15.04  30.1  " 

Repairs 2.00  4.0  " 

Lights,  oil  and  oiling 0.75  1.5  " 

Two  horse-power  at  $0.25  per  day,  per  horse-power. . .  0.50  1.0  " 

Interest  on  cost  of  furnace,  at  6  per  cent,  per  annum. .  0.92  1.84  " 

Totals $24.96  49.94    " 

The  power  required  has  been  determined  by  indicator;  the  fuel, 
from  the  wood  delivered  to  eight  calciners  during  a  mouth;  the 
oil,  lights,  and  proportion  of  labor  in  oiling,  from  the  actual  costs 
at  the  works.  All  these  items,  as  well  as  the  labor  employed  at 
the  furnace,  are  easy  to  arrive  at.  Also,  the  first  cost  of  a  furnace, 
which  can  be  checked  in  various  ways. 

The  only  point  open  to  dispute  is  the  cost  of  repairs.  This  has 
been  taken  from  a  two  years'  run.  The  cost  of  repairs  as  given  by 
H.  C.   Bellinger,  superintendent  Montana  Ore  Purchasing  Com- 


panv,*  Butte,  was  only  II  per  furnace  per  24  hours.  This  figure 
was  arrived  at  from  new  furnaces,  only  six  months  in  operation, 
and  which  had  not  required  many  repairs  nor  new  chains.  A  chain 
costs  about  $130,  and  ou  heavy  sulphides  and  constant  running, 
should,  with  due  skill  and  attention,  last  abont  a  year. 

In  its  construction,  the  9  by  90  foot  Allen-O'Harra  furnace 
requires  125,000  red  brick,  8,000  fire-brick,  36,000  pounds  cast 
iron,  30,000  pounds  wrought  iron,  and  52  perches  stone  work, 
more  or  less. 

Including  excavation,  it  costs  in  Butte  about  16,000. 

The  Pearce  turret  furnace  may  be  described  as  a  long,  narrow 
hearth,  bent  around  a  circle,  the  circumference  of  which  is  a  little 
greater  than  the  length  of  the  hearth,  so  that  the  two  ends  do  not 
quite  meet.  At  this  broken  part  the  roasted  ore  is  discharged. 
The  fresh  ore  is  automatically  fed  from  a  hopper  at  the  other  side 
of  the  break,  and  is  gradually  stirred  and  moved  forward  by  rab- 
bles attached  to  hollow,  air-cooled  arms,  revolving  around  a  sta- 
tionary, central  columo.  The  wall  of  the  hearth  forming  the 
inner  circle  is  provided  with  a  continuous  slot  for  the  sweeping 
passage  of  the  two  revolving  arms,  and  this  slot  is  closed  by  an 
endless  steel  tape,  which  revolves  bodily  with  the  rabble-arms, 
being  continuously  pressed  against  the  slot,  so  as  to  mostly  exclude 
the  cold  air.  The  entrance  of  outside  air  is  still  further  counter- 
acted by  the  employment  of  a  slight  blast  under  the  grate  and 
through  the  hollow  rabble-arms,  which  balances  the  tendency  of 
the  draught  to  suck  air  into  the  furnace,  cools  all  the  exposed  iron 
surfaces,  and  enables  the  metallurgist  to  introduce  an  accurately 
gauged  quantity  of  air,  for  the  purposes  of  combustion  and  oxida- 
tion (900  cubic  feet  per  minute  are  used  at  Argo  when  running 
on  heavy  pyritous  ores).  The  inner  skewback  wall,  that  is  to  say, 
the  wall  immediately  above  the  flue,  is  hung  from  heavy  I-beams, 
whose  extremities  are  supported  by  the  central  column,  and  by  the 
outer  walls  of  the  furnace.  The  bracing  of  the  furnace  is  exceed- 
ingly simple  and  effective,  consisting  merely  of  circular  iron  bands 
for  the  outside,  while  any  distortion  is  prevented  by  radial  struts, 
like  the  spokes  of  a  wheel,  between  the  lintels  and  the  central  col- 
nmn.  Two  or  three  fireplaces  are  spaced  around  the  outer  cir- 
cumference of  the  circle  at  appropriate  points,  the  entering  flame 
being  kept  from  immediate  contact  with  the  ore  by  short  curtain 

*  Engineering  and  Mining  Journal,  July,  22,  1893. 


The  ore  is  stirred  once  in  40  seconds,  or  a  total  of  540  times 
during  the  six  hours  that  it  requires  to  pass  from  feed  to  diseliarge. 
Of  course  the  time  of  roasting  and  number  of  stirrings  can  be 
regulated  to  suit  the  requirements  of  the  material  under  treatment. 
Tbe  greater  length  of  the  outer  circumference  of  the  hearth  as 
compared  with  the  inner  seems  to  have  no  ill  effect  on  the  result, 
the  roasting  being  absolutely  uniform  over  the  entire  width  of  the 
furnace,  and  the  length  of  each  individual  plow-blade  increasing 
slightly  toward  the  outer  circle,  so  that  it  can  move  the  ore  the 
slightly  greater  distance  demanded  by  tbe  increased  size  of  the 
circle.  These  plows  are  simply  plates  of  |-inch  steel,  and  last 
four  to  six  weeks  on  pyrites  containing  40  per  cent,  sulphur.  The 
rabble-arms  that  carry  the  plows  are  of  o-inch  pipe,  and  last  a 
year.  When  the  plows  require  renewal,  the  entire  rabble-arm  is 
uncoupled  outside  of  the  slot,  and  withdrawn  throagh  a  door  in 
the  outer  wall,  a  fresh  one  with  plows  already  in  position  being  at 
once  substituted. 

The  width  of  the  hearth  in  the  original  furnaces  is  6  feet,  but 
some  are  now  being  built  7  feet  wide.  The  diameter  of  the  en- 
closed circular  space  is  19|  feet,  and  of  the  furnace  over  all,  36 
feet.  The  fireplaces  project  6  feet  further,  and  the  entire  furnace 
can  stand  in  a  quadrangle  36  by  42  feet,  thus  occupying  1,512 
square  feet. 

Plates  III.,  IV.,  and  V.  (Figs.  1  to  9),  illustrate  the  Pearce 
turret  furnace.  A  is  the  hearth,  forming  a  circle  with  a  wedge- 
shaped  piece  removed  at  B,  for  the  discharge  of  the  roasted  ore. 
This  hearth  is  constructed  over  the  dust-chamber  C,  through 
which  the  gases  pass  in  a  direction  contrary  to  that  in  which  they 
move  upon  the  hearth.  D  is  the  first  fireplace  and  E  the  second 
one,  the  gases  moving  around  the  hearth  to  the  flue  and  dowutake 
F,  through  which  they  pass  to  the  dust-chamber.  The  inner 
hearth-wall  has  a  continuous  slot  G  (Figs.  3,  4,  5)  for  the  passage 
of  the  spoke-like  rabble-arms  H,  which  have  their  hub  J  around 
the  central  column  I.  This  column  is  stationary,  and  is  hollow 
to  admit  of  the  passage  of  a  light  blast  of  air  to  the  wind-box 
(hub)  .J.  The  superior  portion  of  the  inner  wall  and  skewback 
cannot  be  built  up  in  the  usual  manner,  and  is  therefore  hung 
from  the  eight  12-inch  I-beams  K  by  means  of  stirrups  k^  and  the 
cross-beams  L.  The  rabble-arms  H  are  strongly  braced  by  means 
of  the  straining  rods  7^,  and  are  revolved  by  the  pinion  M  which 
meshes  into  tbe  bull-wheel  N.     This  wheel  is  centered   by  the 

Ill  3TAJ«a 

AV«^(»*  it-a> 

.f^3V440JA.O    TSF. 











rollers  n,  and  the  entire  weight  of  the  rabble-arms  and  driving 
gear  is  taken  by  the  conical  rollers  0  running  on  the  circular  track 
o;  no  weight  at  all  comes  upon  the  hub  J.  A  5-inch  pipe  P  pro- 
tects the  driving-shaft  p  where  it  traverses  the  dust-chamber. 
The  rabble-arms  have  a  joint  at  Q  so  that  they  can  be  adjusted  to 
suit  the  wearing  of  the  plovv-blades.  The  blast  coming  through 
the  pipe  E,  the  central  column  I,  and  the  wind-box  J,  continues 
through  the  rabble-arms  H  (which  consist  of  o-inch  gas-pipe), 
and,  cooling  that  portion  of  the  arms  which  is  exposed  to  the  heat 
of  the  gases,  streams  out  into  the  hearth  through  the  openings  h' 
and  the  little  j)ipes  h"  (Figs.  6  and  7),  thus  cooling  the  plows  and 
furnishing  hot  air  for  the  oxidation  of  the  ore.  On  the  first  por- 
tion of  the  hearth  where  the  fresh  ore  is  being  gradually  heated, 
no  air  is  desired.  The  blast  is,  therefore,  cut  on  opposite  the  ore 
hopper  by  means  of  the  butterfly  valves  aa  (Figs.  1  and  2),  which 
are  closed  by  the  stop  5,  and  again  opened  at  c.  Heated  air  is  also 
introduceri  tljrough  the  exterior  wall  of  the  hearth  by  means  of 
the  intramural  passages  d. 

The  ore  is  dropped  UTaoc  the  hearth  from  the  hopper  S  by  the 
automatic  feed  mechanism  shown  ih  Fig:  ^  and  actuated  by  the 
stops  e  on  the  rabble-arms.  It  is  gradually  advanced  by  the  plows 
in  a  direction  opposite  to  the  gases,  until  it  iis  discharged  at  B  into 
a  car.  The  12-iuch  I-beams  K  take  their  bearings  on  the  central 
column  and  on  the  main  outside  wall  of  the  furnace.  This  cal- 
ciner  is  strongly  banded  externally,  and  is  internally  braced  by  the 
6-inch  struts  T  that  radiate  from  the  central  column.  The  slot  G 
is  closed  by  a  12-inch  steel  tape  U  that  revolves  with  the  rabble- 
arms,  and  is  supported  and  pressed  outward  against  the  walls  of 
the  slot  by  means  of  the  bell-cranks  and  weights  ?/,  Figs.  8  and  0. 
Tlie  bell-cranks  are  supported  on  a  circular  angle  iron  V  that  is 
bolted  to  the  rabble-arms.  The  fire-boxes  burn  slack  coal,  and  are 
provided  with  a  step  grate  W,  Figs.  3  and  5,  and  automatic  coal 
hoppers  X,  Figs.  1  and  2.  The  fireplace  E,  nearest  the  feed,  is 
proviilcd  with  a  curtain  arch  Y,  Fig.  3,  as  the  ore  is  easily  fusible 
at  this  stage  of  the  roasting.  There  are  four  rabble-arras,  but  it 
is  found  best  to  use  only  two  of  them.  The  discharge  vault  is 
provided  with  a  light  stack  Z  to  carry  ofiE  the  fumes. 

About  two  horse-power  are  required  to  run  the  furnace  and 
blast.  Apart  from  repairs  and  renewals,  whicii  are  slight,  no  labor 
is  required  at  tiie  furnace  except  to  oil  the  machinery,  to  fire,  and 
to  have  a  general  supervision  of  its  behavior. 


Some  of  the  results  obtained  in  ordinary  work  by  this  furnace 
are  as  follows: 

Of  iron  pyrites  containing  43  per  cent,  sulphur  and  crushed  to 
pass  a  two-mesh  screen  (9  mm.  openings),  16  tons  per  24  hours  are 
roasted  to  6  or  7  per  cent,  sulphur,  using  2|  tons  of  Colorado 
slack  coal. 

Of  matte  from  the  lead  smelters,  containing  11  per  cent,  lead, 
15  per  cent,  copper,  and  IT  per  cent,  sulphur,  crushed  through  a 
six-mesh  screen  (3  mm.  openings),  11  tons  are  roasted  in  24  hours 
to  3.3  per  cent,  sulphur. 

Of  concentrated  stamp-mill  tailings  (pyrites),  with  45  per  cent, 
sulphur,  and  10  per  cent,  silica,  9  tons  were  dead-roasted  in  24 
hours,  to  show  the  utility  of  the  furmace  for  roasting  for  the  ex- 
traction of  gold  by  chlorination.  No  trace  of  sulphur  remained 
in  the  roast. 

Of  Butte  concentrates  from  the  Gagnon  mine,  consisting  of  vari- 
able mixtures  of  pyrites  and  blende,  but  always  high  in  zinc  and 
sulphur,  15  tons  per  24  hours  are  roasted  to  6  or  7  per  cent,  sul- 
phur. The  following  analysis  represents  an  average  sample  of 
these  concentrates: 

Silica 18.2    percent. 

Iron 20.3 

Zinc 14.85 

Copper 11.29 

Sulphur 31  .5:3 

Total 96  17 

The  Colorado  Smelting  and  Mining  Company,  of  Butte,  has 
erected  double-decked  turret-furnaces,  the  upper  hearth  of  which 
is  supported  upon  an  arch  that  takes  its  peripheral  bearing  upon 
the  main  external  wall  of  the  furnace,  while  its  inner  skewback  is 
supported  by  the  same  interior  wall  that  has  been  already  described 
as  hanging  from  the  heavy  radial  12-inch  I-beams.  The  inner  wall 
above  the  slot  of  the  upper  hearth  being,  in  its  turn,  hung  from  a 
second  set  of  I-beams,  6  feet  higher  than  the  set  belonging  to  the 
lower  hearth. 

This  double  furnace  has  been  only  a  short  time  in  operation, 
but  excellent  results  are  reported  therefrom,  especially  as  regards 
the  consumption  of  fuel.  Each  hearth  is  provided  with  two  fire- 
places, and  Mr.  H.  Williams,  the  manager,  reports  that  while  the 
capacity  for  ore  is  increased,  as  might  have  been  expected,  from 


80  per  cent,  to  100  per  ceufc.,  the  consuiiiptiou  of  fuel  is  only 
heighteued  about  33  per  cent.  This  is  an  extraordinary  and,  to 
me,  unaccountable  saving  in  fuel,  which  I  can  only  explain  by 
assuming  that  much  heat  is  wasted  in  the  single-hearth  furnace; 
probably  because  a  somewhat  high  heat  is  used  just  before  the  end 
of  the  operation,  to  partially  decompose  the  sulphates  remaining 
in  the  roast,  and  much  of  it  must  be  lost,  owing  to  the  short  dis- 
tance between  the  third  fireplace  and  the  stack. 

Indeed,  as  since  the  introduction  of  satisfactory  automatic  cal- 
cining furnaces,  fuel  has  become  the  main  expense  in  the  operation 
of  roasting,  it  seems  a  mistake  that  no  more  is  attempted  in  the 
utilization  of  the  heat  generated  by  the  oxidation  of  the  pyrites. 
When  we  reflect  that  the  heat  thus  produced  is  ample  to  smelt  the 
sulphides  themselves,  as  well  as  an  equal  weight  of  dry  ores,  and 
that  it  is  thus  utilized  in  pyritic  smelting,  we  cannot  fail  to  be 
struck  by  the  seeming  extravagance  of  employing  large  quantities 
of  expensive,  carbonaceous  fuel,  to  burn  up  Nature's  own  fuel  in 
the  ore.  The  actual  quantity  of  heat  generated  by  the  oxidation 
of  sulphides  is  exactly  the  same,  whether  this  oxidation  be  eifected 
in  the  pyritic  smeltiug-furuace,  or  in  the  calciner.  But  in  the 
smelting-furuace,  it  must  be  oxidized  rapidly  in  order  to  generate 
the  intense  heat  necessary  for  fusion,  while  in  the  calciner  the 
oxidation  is  slow  and  quiet,  being  spread  over  several  hours,  so  as 
to  produce  only  the  moderate  temperature  suitable  for  the  process. 
Most  of  this  heat  escapes  through  the  stack  and  in  heating  the 
air  that  is  admitted,  or  finds  its  way,  into  calcining-furnaces.  The 
two  most  obvious  means  of  utilizing  this  slowly-generated  heat, 

1.  By  building  the  hearths  in  such  close  juxtaposition  that  the 
enormous  loss  of  radiation  is  lessened,  and  the  waste  heat  is  stored 
up  in  the  great  n)asses  of  brick-work  forming  the  furnace.  Ex- 
amples: The  improved  Spence  at  the  Parrot  smelter  at  Eutte, 
and  Steinbeck's  multiple-hearth,  circular,  automatic  calciner  at 
Mansfeld,  the  latter  of  which  runs  regularly  on  argentiferous  white 
metal  for  the  Ziervogel  process,  absolutely  without  fuel.  The 
Parrot  furnace  also  runs  for  days  on  heavy  sulphide  ores,  at  the 
rate  of  30  tons  per  day  or  more,  with  cold  fireplaces;  and  when 
fuel  is  used,  it  is  simply  to  increase  the  capacity  of  the  furnace. 

This  type  of  furnace  must  not  be  confounded  with  furnaces  that 
have  their  hearths  built  one  above  another  in  Avhat  a})pears  to  be 
the  same  fashion,  but  Avhere  the  constructors  have  taken  elaborate 


measures  to  carefully  isolate  and  cool  each  individual  hearth.  In 
order  to  save  the  possible  racking  and  distortion  of  the  furnace, 
they  sacrifice  the  main  advantage  of  this  method  of  construction, 
i   0.,  the  conservation  of  the  lieat. 

i  By  eniploving  the  heat  of  calcination  to  preheat  all  air  that  is 
to  enter  either  the  hearth  or  the  ash-pit.  Pearce  pursues  this  plan, 
to  a  certain  extent,  in  his  turret-furnace,  much  of  the  air  entering 
the  hearth  being  preheated  by  its  passage  through  the  rabble-anna, 
or  bv  passing  through  canals  in  the  walls  of  the  furnace.  Bhtke 
carries  this  still  further  in  his  revolving-hearth  Cornish  calciuer 
at  Shullsburg,  Wisconsin,  preheating  the  air  with  the  aid  of  ex- 
traneous carbonaceous  fuel.  Ho  claims  valuable  results  from  this 
svstem,  though  it  seems  a  pity  to  waste  coal  on  preheating  the  air 
when  such  a  vast  store  of  heat  is  available  from  the  operation 

Of  all  the  mistaken  ideas  in  the  construction  of  calciners,  that 
of  cooling  the  hearths,  except  for  the  purpose  of  preheating  the 
air  used  for  this  purpose,  seems  to  me  the  most  illogical.  The 
occasional  disadvantages  of  distortion  can  be  better  borne  than  the 
constant  waste  of  fuel.  It  is  like  cooling  the  hearth  of  a  rever- 
beratory  smelter  by  a  water-jacket,  or  by  the  active  circulation  of 
air  under  a  thin  hearth,  and  then  wondering  why  the  charges  take 
so  long  to  bring,  or  why  they  stick  so  persistently  to  the  bottom. 
As  it  is  now  the  fashion  to  invent  automatic  calciners,  and  as  the 
main  opportunity  in  improvement  lies  in  the  lessening  of  the  fuel- 
consumption,  it  would  be  most  profitable  for  all  aspirants  in  this 
direction  to  spend  a  week  in  working  at  a  battery  of  the  pyrites- 
burners  or  kilns,  as  used  in  the  great  sulphuric  acid  works.  They 
would  at  least  learn  that  the  glowing  brick-work  of  the  burners  is 
the  one  kindler,  regulator,  safety-valve,  and  balance  wheel  of  the 
whole  operation. 

The  tendency  at  present  is  to  drive  calcining-furuaces  rapidly 
and  burn  the  sulphur  and  iron  at  the  highest  allowable  tempera- 
ture by  means  of  the  heat  derived  from  extraneous  fuel,  in  order 
to  obtain  the  greatest  possible  output  from  a  limited  calcining 

Investment  in  plant,  within  reasonable  limits,  is  cheaper  than 
coal  at  $3  to  *(5  per  ton,  and  it  seems  probable  that  slower  run- 
ning, lower  heat  at  the  commencement,  and  through  the  greater 
part  of  the  calcining  process,  and  a  greater  area  of  hearth  per  ton 
of  material  roasted,  will  admit  of  a  more  thorough  utilization  of 


the  heat  evolved  iu  the  combustion  of  the  ore,  and  a  marked  saving 
iu  carbonaceoiis  fuel. 

Both  at  Denver  and  Pueblo,  and  as  well  in  the  O'Harra  as  in 
the  Pearce  calciuers,  the  residues  from  the  distillation  of  the 
Florence  petroleum  are  used  for  firing  to  a  certain  extent,  and  are 
found  a  most  convenient  and  manageable  fuel  for  the  purpose. 

The  costs  of  calcining  in  the  turret  furnace  have  been  looked 
into  and  discussed  by  many  of  our  copper  men,  as  these  furnaces 
are  in  regular  operation  at  three  of  the  greatest  smelting  centers 
iu  the  country:  Denver,  Pueblo,  and  Butte.  But  Mr.  Pearce  of 
Argo  has  kindly  furnished  me  with  some  exact  figures  from  the 
Argo  records  that  are  of  value  to  the  profession. 

Three  turret  furnaces  were  run  on  a  certain  pyritic  ore  from 
December  2,  1893,  to  January  20,  1894.  In  this  period  of  48 
days  there  were  calcined  2,319.558  tons  of  ore,  being  a  trifle  more 
than  16.1  tons  per  furnace  per  day.  The  ore  contained  about  25 
per  ceut.  silica  and  75  per  cent,  sulphides,  mostly  pyrite.  Its 
sulphur  contents  averaged  36  per  cent.  It  was  roasted  down  to 
about  4.75  per  cent,  sulphur  at  the  following  cost,  the  transporta- 
tion to  and  from  the  furnace  being  omitted,  as  it  is  a  variable  item 
at  different  smelters,  and  has  nothing  to  do  with  the  cost  of 


Total  Cost 

Expenses.  per  Ton. 

Labor — 1  man  per  12  hour  shift  at  $2.25,  and  extra 

labor $235.60  9.73  cents. 

Coal— 235.47  tons  at  $2  15 $506.26 

132.47        "       1.55 205.33 

367.94  tons  unloading  at  8  cents,    29.43 

741.01  31.95  " 

Repairs — new  rabb'es  and  sundries 16.00  0.69  " 

Power,  steam,  and  oil 180.00  7.76  " 

Interest  on  furnaces  at  6  per  cent,  per  annum 128.10  5.52  " 

Total  cost $1,290.71  55.65      " 

The  following  table  gives  the  cost  of  calcining  pyritic  and  zincky 
concentrates  from  the  Gagnon  mine  at  Butte  in  the  turret  fur- 
naces at  Ai'go.  A  partial  analysis  of  these  concentrates  is  given 
on  a  preceding  page.  This  contained  14.85  per  cent,  zinc  and 
31.53  per  cent,  sulphur,  and  were  roasted  down  to  7.44  per  cent, 
sulphur  at  the  rate  of  16.889  tons  per  day;  1.02  per  cent,  of  the 

212  MODERN    COPPER    SilELTlXG. 

residual  sulphnr  was  present  as  zinc  sulphate;  152.093  tons  were 
calcined  in  nine  days. 


Total  Cost 

Expense.  per  Ton. 

Labor — as  above $14.12  9.38  cents. 

Coal— 17.94.5  tous  at  $2.15 $38.58 

9.595        '•       2.00 19.19 

27.54  tons  unloading  at  8  cents,    2.20 

59.97  39.43 

Repairs — new  rabbles 4.00  2.63 

Power,  steam  and  oil   9.00  5.91 

Interest  on  furnace  at  6  per  cent,  per  annum 8.01  5.26 

Total  cost $95.10  62.51 

At  Butte,  Montana,  with  wages  at  $3.50  per  day,  and  poor  coal 
at  $3.50  per  ton,  the  cost  of  roasting  the  above  concentrates  in  the 
turret-furnaces  is  about  68  cents  per  ton. 

Heavy  Leadville  pyrites,  containing: 

Iron . ...  41  per  cent. 

Sulphur 46 

Silica 5        " 


is  roasted  down  to  4.46  per  cent,  sulphur,  at  the  rate  of  14.768 
tous  per  day,  at  a  cost  of  about  57  cents  per  ton. 

Concentration-matte  from  the  lead  smelters,  containing: 

Copper 34.4  per  cent. 

Iron 18.3 

Sulphur 21.3 

Lead 11.8 


was  roasted  down  to  6.89  per  cent,  snlphnr,  at  the  rate  of  13.010 
tons  per  day  per  furnace;  104.154  tons  of  this  matte  were  roasted 
in  eight  days,  with  the  following  costs: 


Total.  Per  Too. 

Labor,  as  before $12.54  12.04  cento 

Coal— 26.8  tons 53.49  51.35      " 

Repairs — new  rabbles 4.00  3.84      " 

Power,  steam,  and  oil 8.00  7.68      " 

Interest  on  furnace  at  6  per  cent,  per  annum 7.12  6.88      " 

Total  cost $85.15  81.74      " 


The  turret  furnace  would  seem  peculiarly  adapted  to  the  cal- 
cination of  auriferous  pyrites  for  extraction  by  chlorination. 
Indeed,  several  are  now  constructing  for  that  purpose. 

In  a  trial  run  at  Argo,  concentrated  tailings  from  the  stamp- 
mills  of  Gilpin  County,  Colorado,  containing  about  79.5  per  cent, 
pyrite,  representing  42.1  per  cent,  sulphur,  were  roasted  down  to 
0.22  per  cent,  snlphur  at  the  rate  of  9.813  tons  per  furnace  per 
day.  In  8|  days,  83.411  tons  were  calcined,  with  the  following 


Total.  Per  Ton. 

Labor,  as  before $23.69  28.4  cents. 

<Joal— 11.64  tons  at  $2.30 $26.77 

7.981     "        1.75 13.96 

19.621  tons  unloading  at  8  cents,  1.56 

42.29  50.7 

Repairs — new  rabbles,  etc 6.20  7.43 

Power,  steam,  and  oil 7.42  8. 90 

Interest  on  furnace  at  6  per  cent,  per  annum 7.56  9.07 

Total  cost $87.16      $1.04.5 

Fhie-dust.  —  As  may  be  inferred  from  the  quiet  and  regular 
mechanical  movements  that  occur  in  the  turret  furnace,  its  pro- 
duction of  flue-dust  is  very  small.  In  cleaning  up  the  dust- 
chambers  and  flues  after  a  run  of  2,720.542  tons  of  ore,  22.65  tons 
of  dust  were  recovered,  being  0.8  per  cent. 

Tlie  cost  of  a  turret  furnace  at  Argo,  as  built  by  the  inventor, 
Mr.  Pearce,  is  $5,460.70,  inclusive  of  royalties. 

I  am  indebted  to  the  kindness  of  Mr.  A.  S.  Dwight,  superin- 
tendent, for  tlie  cost  of  the  two  new  turret-furnaces  erected  at 
The  Colorado  Smelting  Company's  Works  at  Pueblo.  The  total 
expense,  including  royalties,  was  $12,296,  or  $6,148  each.  In  this 
case  there  were  some  extra  expenses,  owins  to  necessarily  exten- 
sive foundations,  fire-brick  hearths,  arches,  etc. 

The  turret-furnace  is  a  model  calciner  in  its  running,  and  in 
the  manner  in  which  its  mechanical  details  have  been  worked  out. 
It  is  entirely  automatic  in  its  action,  one  man  attending  three  or  more 
furnaces.  It  requires  but  little  power  to  run,  and  its  repair-bill  is 
mainly  confined  to  changing  plow-blades  once  in  four  to  six  weeks, 
and  in  renewing  rabble-arms  annually.  In  only  one  respect  does 
it  seem  to  me  open  to  criticism,  and  that  is  in  its  consumption  of 


fuel.  This  is,  on  good,  pyritic  ores,  some  lU  per  cent,  to  18  per 
cent,  of  the  weight  of  the  ore;  and  though  it  must  be  remembered 
that  Argo  conditions  demand  a  considerably  more  thorough  calci- 
nation than  is  required  at  Butte,  and  that  it  takes  more  fuel  to 
reduce  the  sulphur  in  an  ore  from  10  per  cent,  down  to  5  per  cent, 
than  to  lower  it  from  40  per  cent,  to  25  per  cent.,  yet  there  is, 
nevertheless,  too  great  a  loss  of  heat,  and  too  little  use  made  of 
the  caloric  generated  by  the  oxidation  of  the  sulphur  and  iron  in 
the  furnace. 

That  this  is  the  principal  direction  in  which  we  must  look  for  a 
still  greater  reduction  in  the  cost  of  calcination  is  evident,  when 
we  note  that  the  fuel,  even  at  the  comparatively  low  price  of  coal 
in  Colorado,  forms  about  60  per  cent,  of  the  total  cost  of  roasting 
pyritic  ores  down  to  from  4  per  cent,  to  7  per  cent,  sulphur. 

Tlie  imjjroved  Spence  calcining  furnace  was  designed  and 
erected  for  the  Parrot  Silver  and  Copper  Company,  by  Messrs. 
Keller,  Gaylord,  &  Cole.  The  company  has  lately  added  two  new 
ones,  and  now  has  three  of  them  running  at  its  smelter  at  Butte, 
Montana,  these  having  displaced  the  twelve  lougreverberatory  cal- 
ciners  there  in  use,  as  well  as  the  ordinary  Spence  furnaces  which 
were  erected  at  the  Parrot  some  three  years  ago.  The  improved 
Spence  was  originally  designed  as  a  circular  furnace,  though  the 
stirring  arms  returned  idle  on  their  track, without  ever  completing 
the  entire  revolution,  as  in  the  other  circular  calciners.  But  the  in- 
ventors eventually  settled  on  the  present  rectangular  form,  and 
the  furnace  is  now  built  as  two  sets  of  five  hearths  and  a  drying- 
hearth,  the  driving  mechanism  being  between  these  two  blocks, 
and  the  whole  structure  constituting  a  single  furnace. 

There  are,  of  course,  six  sets  of  rabble-arms  on  each  side,  one 
set  above  the  other,  projecting  through  slots  into  their  respective 
hearths.  The  rabble-arms  are  provided  with  plows  both  above 
and  below,  as  in  the  O'Harra  furnace,  and  these  plows  are  only  in 
contact  with  the  ore  when  traveling  in  one  direction.  When  their 
motion  is  reversed,  a  tripping  mechanism  turns  the  arm  one- 
fourth  of  a  revolution,  so  that  both  its  sets  of  plow-blades  lie  hori- 
zontally above  the  ore,  and  in  this  position  the  rabbles  move  back 
to  the  other  end  of  the  furnace.  When  they  reach  this  point,  the 
arm  is  again  tripped  and  revolves  90  degrees.  But  the  revolution 
of  the  arm  always  continues  in  the  same  direction,  so  that  the 
plows  that  were  at  first  projecting  perpendicularly  into  the  air  are 
now  brought  into  use.     By  this  ingenious  device  the  plows  arc 


.  e 

—     -4- 





[—1        1 
I    ,    i- 





—  —  —  +-  T  *-"C"- 




Keli  er-Gatlchd-Coi. 

Sectional  Plan. 


•,^oiT0s8  eeoft') 

Keller-Gatlord-Cole  Calciner, 


euabled  to  cool  oflt',  and  the  two  sets  of  plow-blades  are  so  fitted  on 
the  rabble  that  they  constantly  alternate  in  the  ridges  and  furrows 
of  the  ore  on  the  hearth. 

The  driving  gear  consists  of  a  wire  rope,  the  extremities  of 
which  are  attached  to  the  rabble-frames,  while  the  ropes  them- 
selves pass  around  a  large  driving-wheel,  on  whose  shaft  is  keyed 
a  pinion  that  receives  reciprocal  motion  from  a  rack  actuated  by 
a  hydraulic  piston. 

The  slots  are  closed  by  traveling  steel  tapes,  as  in  the  turret 
calciner;  but  this  furnace  being  longitudinal,  and  the  motion  of 
the  rabbles  being  reciprocally  to  and  fro,  the  tapes  are  wound  and 
unwound  alternately  on  horizontal  pulleys,  placed  at  each  end  of 
the  hearth.  These  are  governed  by  springs  so  as  to  keep  the  tape 
taut,  and  its  winding  is  assisted  by  counter  weights. 

The  hearths  are  three  feet  ai^art  vertically,  and  are  covered  with 
siliceous  tailings  from  the  concentrator.  The  enormous  mass  of 
brick-work  contained  in  the  superincumbent  hearths  and  arches 
retains  much  of  the  heat  generated  by  the  oxidation  of  the  sul- 
phides, and  consequently  diminishes  the  fuel  consumption  to  a 
point  that  would  seem  impossible  to  those  who  have  not  given 
attention  to  this  particular  subject. 

There  is  a  2|  by  4  foot  fireplace,  fired  with  slack  coal,  to  each 
block  of  hearths;  that  is  to  say,  two  fireplaces  to  the  double  block 
forming  a  single  furnace.  The  flame  is  only  allowed  to  traverse 
the  top  hearth,  where  it  is  nsed  to  ignite  the  sulphur  quickly, 
the  temperature  on  the  lower  hearths  being  ample  without  extra- 
neous heat  to  reduce  the  sulphur  to  the  required  standard  7  to  10 
per  cent.  I  am  informed  that  by  using  more  time  and  fuel,  there 
has  been  no  difficulty  experienced  in  reducing  the  sulphur  to  any 
desired  limit. 

The  following  results  are  taken  mainly  from  written  statements 
made  to  me  by  Mr.  H.  A.  Keller,  superintendent  of  the  Parrot 
smelter  and  one  of  the  inventors  of  the  furnace,  and  therefore 
cannot  carry  the  same  weight  as  though  made  by  unprejudiced 
observers.  But  it  is  only  just  to  say,  that  personal  observation 
and  careful  questioning  of  the  workmen  employed  about  the 
smelter,  especially  in  regard  to  repairs,  stoppages,  and  fuel  con- 
sumption, have  failed  to  detect  any  exaggeration  in  the  claims 

The  furnace  has  been  mainly  run  on  mixed  sizes  of  concevvtrates 
from  the  Parrot  mine,  of  which  the  following  was  the  average 
composition  for  the  first  nine  montlis  of  1891: 


Copper 9. 8  per  cent. 

Iron 33.8 

Silica    13.3 

Sulphur 41.2         " 

Silver 0.027     "        (8  oz.  per  ton). 

98.127     •* 

Mr.  Keller  states  that  while  roasting  45  tons  (90,000  poncds) 
per  24:  hours  of  the  above  concentrates,  the  farnace  has  during  the 
past  three  mouths  burned  three-fourths  of  a  ton  of  slack  coal  (at 
#3.50  per  ton).     The  coal  averages  about  IS  per  cent.  ash. 

The  ore  is  fed  to  the  calciuer  automatically  by  heavy  fluted 
rollers;  and  as  the  bringing  of  the  raw  ore  to  the  furnace,  and  the 
removal  of  the  calcined  ore  depend  for  their  cost  upon  the  general 
arrangement  of  the  plant,  and  are,  therefore,  so  variable  at  differ- 
ent works  as  to  completely  invalidate  any  exact  inquiry  into  the 
comparative  cost  of  roasting  in  different  types  of  calciuers,  I  have 
entirely  omitted  them  in  every  case,  preferring  to  let  each  smelter 
calculate  the  cost  of  the  above  items  to  suit  his  individual 

Since  the  reverberatory  calciners  have  been  given  up  at  the 
Parrot  smelter  there  is  no  roasting  foreman.  The  three  improved 
Spence  calciuers  are  attended  by  one  man  per  l"2-hour  shift,  who 
fires  (handling  f  ton  coal  for  each  furnace),  and  beyond  this  sim- 
ply has  to  oil  and  oversee  the  machinery.  As  his  wages  are  84  per 
shift,  and  the  amount  of  ore  handled  per  shift  by  the  three  cal- 
ciners is  6T^  tons,  the  Cost  of  labor  per  ton  is  not  quite  6  cents. 

Using  f  ton  coal  per  shift,  at  ?3.50  per  ton,  and  roasting  22^  tons 
of  ore,  the  cost  for  fuel  per  ton  of  ore  is  5.83  cents.  The  furnace 
has  been  run  for  several  successive  days  without  any  fuel  at  all, 
the  duty  being  reduced  from  45  to  30  tons  ore  per  24  hours. 

It  is  stated  to  require  two  horse-power  to  run  the  furnace. 

I  find  from  personal  inquiry  that  most  of  the  Butte  metallurgists 
who  have  carefullv  followed  the  development  and  operation  of  this 
furnace  seem  inclined  to  admit  the  correctness  of  the  above  state- 
ments so  far  as  regards  labor  and  fuel  consumption,  but  are  not  in 
a  position  to  express  a  positive  opinion  as  to  the  repairs. 

I  examined  the  record  of  the  furnace  on  the  Parrot  books  and 
found  that  its  stoppages  were  about  12  hours  per  month,  mainly 
for  renewing  rabble-arms  and  attending  to  the  steel  tape  that 
closes  the  slots. 


Mr.  Keller's  own  statements  (December  2.  1894),  regarding  the 
total  repairs  on  one  furnace  for  the  past  12  mouths,  are  as  follows: 

36  sets  of  plow-blades  at  $8.94 , $331.84 

1  fall  set  of  4-inch  pipes  for  arms  (12  pipes,   each  7  feet 

long) 29.40 

Other  repairs,  averaging  $5  per  month 60.00 

Total $411.24 

being  about  $1.13  per  day.  Mr.  Keller  calls  the  repairs  $1.25  per 
day,  or  2.78  cents  per  ton  of  ore.  The  rake-end  is  the  only  portion 
of  the  rabble-arm  exposed  to  heat,  and  its  life,  when  running  45 
tons  ore  per  day,  is  four  months  more  or  less,  according  to  whether 
it  belongs  to  one  of  the  hotter,  or  one  of  the  cooler  hearths.  As 
they  form  the  main  item  of  repairs,  it  is  interesting  to  know  their 
cost  in  detail. 


7  feet  4-inch  pipe  at  35  cents , $2.45 

18  cast-iron  plow-blades,  7  pounds  each,  at  4  cents 5.04 

19  six-inch  bolts  at  10  cents 1.90 

One-half  day  machine  work  at  $4 2.00 

Total $11.39 

It  is  claimed  by  the  inventors,  that  there  is  now  no  racking  of 
the  furnace,  nor  distortion  of  slot.  There  are  no  fire-brick  used 
in  the  furnace,  except  where  red  brick  are  so  fusible  as  to  be  unfit 
for  lining  the  fire-box. 

It  will  be  interesting  to  assemble  the  figures  already  given,  and 
thus  determine  the  cost  of  roasting  at  the  Parrot  smelter,  as 
claimed  by  Mr.  Keller  and  his  associates. 

The  cost  of  erecting  one  of  these  45-ton  improved  Spence  fur- 
naces at  Butte  is  about  §10,000.  The  interest  on  the  above  sum, 
at  6  per  cent,  per  annum,  would  amount  to  3.G  cents  per  ton  of  ore. 

COST     OF    ROASTING     ONE    TON   (2,000    POUNDS)     ORE     IN     IMPROVED    SPENCE 


These  figures  are  deduced  from  H.  A.  Keller's  statements,  based  on  twelve 
months'  running  (transportation  of  ore  to  and  from  furnace  is  not  included). 

Labor — per  ton  of  ore 6.00  cents. 

Fuel  "  "       5.83     " 

Repairs       "  "     2.78     " 

Power  and  oil,  per  ton  of  ore 2.22     " 

Interest  on  copt  of  furnace  per  ton  of  ore 3.06     " 

Total - 20.43     " 

Or  about  20^  cents  per  ton  of  raw  ore. 


While  these  unusually  low  figures  are  based  primarily  on  Mr. 
Keller's  own  figures  at  the  Parrot  smelter,  I  should  not  publish 
them  did  I  not  believe  them  to  be,  in  the  main,  correct.  But  a 
personal  examination  of  the  furnace,  and  a  recent  visit  to  the 
Mansfeld  works  in  Germany,  where  approximately  identical  results 
have  been  obtained  for  several  successive  years  in  Dr.  Steinbeck's 
modified  Parkes  calciners  (calcining  white  metal  without  fuel,  for 
the  Ziervogel  silver  extraction)  has  enabled  me  to  assimilate  these 
results  with  less  astonishment  than  many  metallurgists  will  prob- 
ably experience.  The  main  doubtful  point  with  me  is  the  question 
of  repairs,  and  on  this  point  I  have  not  had  a  sufficiently  long 
acquaintance  with  the  furnace  to  express  an  intelligent  opinion. 

The  Broion  liorseshoe  furnace  is  also  annular  like  the  turret  fur- 
nace. But  it  is  bent  around  a  larger  circle,  the  diameter  of  the 
unoccupied  space  in  the  center  being  41  feet  10  inches,  and  the 
outer  diameter  68  feet  2  inches.  With  its  external  fireplaces,  it 
occupies  a  quadrangle  of  73  feet,  or  an  area  of  5,329  square  feet. 
The  hearth  proper  is  8  feet  wide  in  the  clear  and  occupies  about 
four-fifths  of  the  circle,  the  remaining  fifth  being  completely  cut 
out,  the  free  space  thus  formed  being  used  to  cool  the  rabbles. 
By  means  of  projecting  tiles  in  roof  and  floor,  a  narrow  gallery  is 
formed  on  either  side  of  the  hearth.  The  gallery  on  the  outer 
circumference  contains  simply  a  rail  of  hard-baked  tile,  on  which 
runs  the  outer  wheel  of  the  stirring  carriage.  The  inner  gallery 
contains  an  iron  rail  for  the  inner  wheel  of  the  same  carriage,  and 
also  the  horizontal,  grooved,  idler-pulleys  which  guide  the  driving 
cable.  This  cable  is  driven  by  a  simplified  adaptation  of  the  means 
employed  on  cable-roads,  consisting  of  a  grip-wheel  with  tightener 
and  guide-sheaves. 

The  cable,  guide-sheaves  and  inner  rail  are  cooled  by  admitting 
a  little  air  around  each  sheave  into  the  inner  gallery,  and  it  is 
undoubtedly  a  valid  claim  of  the  inventor,  that  when  the  furnace 
is  properly  run,  none  of  this  iron  work  becomes  hot  enough  to 
seriously  scorch  the  naked  hand. 

The  ore  is  charged  from  an  ingenious,  automatic  hopper  and 
apron,  and,  as  in  all  similar  calciners,  is  gradually  carried  around 
to  the  other,  or  discharge-end,  by  means  of  plows,  which  are  at- 
tached to  carriages,  running  on  the  two  rails  already  described. 
These  carriages  and  their  attached  plows  are  intermittently  cooled 
in  a  very  peculiar  and  original  manner.  There  is  alM'ays  one  car-  , 
riage  standing  idle  on  the  rails  where  they  cross  the  open  space 

!.■  l]il:M 









'  .^ 



between  the  adjacent  ends  of  the  hearth.  It  requires  about  two 
minutes  for  each  stirrer  to  make  the  circuit  of  the  hearth,  so  that 
the  idle  one  has  this  same  length  of  time  to  cool  off  in.  After  its 
emergence  from  the  hearth,  the  moving  (heated)  carriage  comes 
in  contact  with  the  cooled  one  that  is  at  rest,  pushing  it  forward 
a  short  distance,  until  the  carriage  in  the  lead  becomes  attached 
to  the  driving-cable  by  means  of  an  automatic  gri]^,  the  heated 
carriage  being  detached  at  the  same  moment.  The  Collinsville 
Zinc  Company  of  Illinois,  and  the  Glendale  Zinc  Company  of 
South  8t.  Louis,  Missouri,  report,  after  steadily  running  the  fur- 
nace for  several  months,  that  the  action  of  the  grip,  cable  and 
sheaves  is  satisfactory,  in  spite  of  the  high  temperature  used  in 
roasting  zincbleude. 

The  Brown  horseshoe  calciner,  as  built  by  Fraser  &  Chalmers, 
is  illustrated  on  Plate  VIII. 

The  annular  hearth  A  is  broken  at  B  for  the  ore-discharge,  and 
to  afford  a  cooling  space  for  the  plows.  These  are  not  shown  in 
the  drawing,  but  are  mounted  on  wheels  running  upon  rails  in  the 
lateral  galleries  C  and  D,  Figs.  1  and  2.  The  inner  rail  c  is  of 
iron;  the  outer  one  d,  of  hard-baked  tile,  except  in  the  broken 
portion  of  the  furnace.  The  plow  carriages  are  moved  by  an  end- 
less cable  F,  Fig.  5,  which  runs  around  the  little  horizontal  rollers 
E,  and  is  driven  by  the  ordinary  cable-car  mechanism,  shown  in 
perspective  in  Fig.  5.  The  gases  flame  from  the  three  lire-boxes  G, 
H,  and  I  enters  the  hearth  and  passes  out  through  the  flue  J  into  the 
stack  k.  The  heated  plow  which  has  just  completed  the  circuit 
of  the  furnace,  comes  into  the  open  air  at  L.  It  soon  comes  in 
contact  with  the  cooled  carriage  that  has  been  standing  in  the 
open  for  some  minutes,  and  pushes  it  ahead  to  where  it  is  gripped 
by  the  cable  at  M,  the  heated  carriage  remaining  in  the  place  of 
the  cooled  one.  The  ore  is  fed  from  the  hopper  N,  and  is  dis- 
charged at  H.  The  rollers  E,  which  are  mostly  outside  of  the 
hearth  (see  Fig.  2.),  the  cable  F,  and  the  rail  r,  are  said  to  be  so 
cooled  by  the  external  air  and  inward  draught  as  never  to  reach  a 
temperature  of  150  degrees  Fahr.  (65  degrees  Cent.).  Air  is  also 
admitted  through  the  roof  by  means  of  the  holes  0.  As  there  are 
no  revolving  arms,  the  hearth  is  braced  with  tie-rods  in  the  usual 

It  is  stated  that  1|  horse-power  is  required  to  run  the  machinery. 
Also  that  in  roasting  heavy  ziucblende  ores,  about  20,000  pounds 
of  finished  product  is  made  per  24  hours,  the  ore  averaging  over 


30  per  cent,  solphnr,  and  being  roasted  down  to  0.85  per  cent,  to 
1  per  cent.  There  are  four  fire-boxes  on  the  Collinsville  furnaces, 
and  about  12,000  pounds  of  refuse  slack  from  the  adjacent  coal 
mines  is  used  per  24  hours. 

Of  course,  this  is  no  fair  test  as  to  what  the  furnace  would 
accomplish  on  ordinary  pyritic  ores,  but  there  seems  no  reason  to 
doubt  that  it  will  oxidize  as  rapidly  and  effectively  as  any  of  its 
rivals  under  equal  conditions. 

The  Consolidated  Kansas  City  Smelting  and  Befining  Company 
has  just  erected  one  of  these  furnaces  for  the  sulphate-oxide  calci- 
nation of  its  copper-lead  matte,  for  the  Hunt  &  Douglas  copper- 
extraction  process.  The  company  inform  me  that  they  are 
roasting  mattes  for  the  Hunt  &  Douglas  wet  extraction  process, 
and  containing  10  per  cent,  to  20  per  cent,  lead  and  25  per  cent, 
to  35  per  cent,  copper,  at  the  rate  of  3(3,000  pounds  per  furnace 
per  24  hours,  with  3^  to  3f  tons  slack  coal.  The  calcined  ore 
contains  8T  per  cent,  to  92  per  cent,  of  its  copper  soluble  in  the 
Hunt  &  Douglas  bath. 

Brown  urges,  as  a  valuable  feature  of  his  furnace,  the  long  road 
that  each  particle  of  ore  has  to  travel.  He  claims  that  it  is  thus 
peculiarly  suited  to  the  roasting  of  easily  fusible  ores,  as  they  are 
advanced  so  slowly  and  gradually  toward  the  hotter  portion,  that 
the  sulphides  have  ample  time  to  decompose  and  lose  their  extreme 
fusibility  before  being  subjected  to  a  temperature  higher  than 
they  can  bear. 

The  furnace  has  stood  a  severe  test  in  its  satisfactory  work  on 
zincblende  ores  for  more  than  a  year,  and  has  now  entered  into 
competition  with  the  other  automatic  copper-calciners. 

TJie  Spence  automafic  desiilphurizer  is  a  Maletra  furnace  im- 
proved and  provided  with  automatic  rakes:  It  is  extensively  used 
and  is  too  well-known  to  require  a  detailed  description.  Fig.  25 
gives  a  longitudinal  section  in  detail.  It  is  used  much  for  roasting 
fines  for  sulphuric  acid  manufacture,  but,  in  its  present  form,  has 
too  small  a  capacity,  and  requires  too  much  power  per  ton  of  prod- 
uct to  compete  with  the  newer  automatic  calciners.  At  the  Par- 
rot smelter  in  Butte,  Montana,  a  double  Spence  furnace  has 
roasted  16,000  pounds  of  concentrates  per  24  hours,  reducing  the 
sulphur  from  40  per  cent,  to  8  per  cent.  This  is  an  unusual  duty, 
and  vet  is  much  too  small  for  prevailing  conditions.  The  cost  of 
roasting  at  the  Parrot,  in  these  furnaces,  is  reported  to  me  to  be 
about  81.25  per  ton  of  ore. 



The  accompanying  drawiug  shows  a  Hammond  improved  Spence 
furnace  used  at  the  great  Tread  well  mill,  Douglas  Island,  Alaska, 
where  a  number  of  them  are  employed  in  roasting  the  gold-bearing 
concentrates  for  treatment  by  the  Plattner  chlorination  process; 
six  double  furnaces  roast  from  18  to  20  tons  of  concentrates  a  day 
to  a  "dead  roast,"  with  an  expenditure  of  about  one-eighth  cord 
of  wood  per  ton  of  ore.  The  space  required  is  small  and  no  skilled 
labor  is  necessary.  Once  adjusted,  it  will  continually  discharge  a 
finished  product.  Two  men  on  a  shift  can  attend  to  six  double 
furnaces  easily.  One  keeps  the  hoppers  full  while  the  other  keeps 
the  temperature  even.  The  fronts  and  backs  of  the  furnaces  are 
so  arranged  that  the  supply  of  ore  can  be  regulated  exactly.  The 
dust  is  even  less  than  in  the  old  reverberatory.  A  substantial 
hydraulic  cylinder  moves  the  rakes,  which  are  so  arranged  as  to 
prevent  the  banking  of  the  material  at  the  ends  of  the  furnace. 
The  iron  rails  of  the  Spence  furnace,  which  gave  much  trouble, 
are  replaced  in  this  by  very  hard  brick  tiles.  With  ordinary  care 
the  iron  rakes  will  last  six  months  when  salt  is  used  in  roasting, 
and  two  years  when  it  is  not  employed,  and  when  burnt  out  can 
be  replaced  by  new  rakes  in  ten  minutes. 

Hie  Math  lessen  d-  HegeJer  Zinc  Company  of  La  Salle,  Illinois, 
has  developed  since  1889  a  peculiar,  but  for  zinc  ores  effective 
type  of  calcining  furnace.  In  estimating  its  work,  it  must  be 
remembered  that  it  is  used  solely  for  the  sweet-roasting  of  zinc- 
blende  ores,  and  that  its  gases  are  employed  for  the  manufacture 
of  sulphuric  acid,  when  it  is  so  desired. 

It  consists  of  two  seven-storied  hearths,  built  side  by  side  in  one 
block,  the  hearths  being  4-^  by  46  feet,  and  possessing  a  common 
division  wall.  The  furnace  is  heated  by  generator  gas,  the  flame 
passing  back  and  forth  under  the  three  lower  hearths,  the  upper 
ones  receiving  no  extraneous  heat.  There  is  one  rake  for  each 
double  hearth,  and  this  implement  rests  most  of  the  time  on  a 
swinging  platform  at  the  end  of  the  hearth.  About  once  an  hour, 
the  rake  is  attached  by  hand  to  an  iron  bar  that  is  pushed  through 
the  hearth  from  the  opposite  end,  and  is  then  dragged  back 
through  the  ore,  the  bar  being  moved  by  friction  pulleys.  The 
outside  platform,  on  which  the  rake  normally  rests,  can  be  swung 
around  opposite  the  opening  of  the  twin  hearth,  and  is  then  drag- 
ged back  through  the  latter  in  the  same  manner  being  thus 
exposed  but  a  short  time  to  the  high  temperature  of  the  hearth. 
The  company  inform   me  that  a   double,  seven-story  furnace  pro- 


duces  40,000  pounds  of  thorouglily  roasted  ore  per  24  hours,  from 
7,000  pounds  of  zincblende,  with  a  consumption  of  9,600  pounds 
of  refuse  slack  coal. 


Blake*  describes  a  tabular,  revolving  roaster  with  automatic  feed 
and  delivery,  that  is  said  to  be  an  improvement  on  Bruutou's 
Cornish  calcining  furnace.  It  is  intended  and  used  for  calcining 
the  iron  pyrites  in  the  impure  zincblende  of  Shullsburg,  Wiscon- 
sin, so  that  it  may  be  easily  removed  from  the  blende  by  mechanical 
concentration.  It  oould,  of  course,  bo  adapted  for  copper  ores. 
It  consists  of  a  circular,  terraced  table,  16  feet  in  diameter,  cov- 
ered with  fire-brick,  and  made  to  revolve  slowly  (10  revolutions  per 
hour)  in  a  horizontal  plane.  It  is  supported  upon  cast-iron  balls 
running  in  a  grooved,  circular  track  12  feet  in  diameter,  and  is 
covered  with  a  dome-shaped  arch.  Plows  fixed  in  the  roof  stir 
the  ore,  and  gradually  urge  it  downhill  toward  the  circumference 
of  the  hearth.  Careful  arrangements  are  made  for  the  introduc- 
tion of  pure  air,  strongly  preheated  by  two  Siemens'  accumulators. 
As  no  assays  or  analyses  are  as  yet  made  public,  and  as  the  purpose 
and  conditions  of  the  calcination  at  Shullsburg  are  totally  different 
from  the  requirements  of  the  copper  metallurgist,  it  is  impossible 
to  institute  any  comparisons  as  to  results.  In  calcining  the  pyrite 
in  a  mixture  consisting  of  equal  parts  of  pyrite  and  zincblende  in 
wheat-sized  grains,  Mr.  Blake  states  that  20  tons  per  24  hours  is 
the  regular  duty  of  a  16-foot  furnace. 

*  Transactions  American  Institute  Mining  Engineers,  Vol.  XXL,  p.  943. 



The  object  of  smelting  ores  of  copper  is  to  effect  a  separation  of 
the  metal  by  a  mecbauical  process  of  concentration,  many  chemical 
changes  important  to  the  result  also  occurring  before  the  worthless 
and  valnable  portions  of  the  ore  can  separate  according  to  their 
specific  gravity.  The  entire  mass  of  rock  which  contains  the  cop- 
per (often  also  gold  and  silver)  must  be  rendered  so  liquid  that 
the  metallic  or  sulphide  portions  can  freely  sink  to  the  bottom, 
whence  they  can  be  drawn  ofE  separately,  while  the  worthless 
molten  rock  (slag)  floats  on  the  surface,  and  is  reimoved  by  appro- 
priate means. 

In  smelting  sulphide  ores,  we  cannot  profitably  produce  metallic 
copper  at  a  single  operation ;  for  both  the  cost  of  removing  all  the 
sulphur  (calcination),  and  the  tenor  of  the  slag  would  be  too  high. 
The  greater  portion  of  the  sulphur  is  removed  from  the  ore  by  cal- 
cination, and  the  remaining  sulphur  combiues  with  the  copper, 
and  with  a  certain  amount  of  iron,  to  form  the  matte  or  regulus 
which  is  the  object  of  our  exertions,  and  which  may  be  regarded 
as  a  highly  concentrated  ore,  free  from  gangue  rock  and  containing 
90  per  cent,  of  the  copper,  90  per  cent,  of  the  silver,  and  99  per 
cent,  of  the  gold  that  was  present,  hy  assay,  in  the  original  ore. 
(Of  course  these  results  vary  considerably,  according  to  degree  ox 
concentration,  composition  of  ores,  etc.) 

It  will  be  at  once  apparent,  that  the  higher  the  degree  of  con- 
centration, i.e.,  the  more  tons  of  ore  we  can  put  into  one  ton  of 
matte,  the  lighter  will  be  the  future  cost  of  refining  this  matte, 
per  tun  of  original  or«.  For  instance:  if,  in  smelting  12  tons  of 
ore,  we  can  throw  11  tons  over  the  dump  in  the  shape  of  slag,  and 
concentrate  the  entire  value  of  the  12  tons  of  ore  into  one  ton  of 
matte,  the  cost  of  refining  that  matte,  at  §18  per  ton,  will  be 
divided  by  twelve,  thus  being  only  $1  ."30  on  the  ton  of  original 
ore.     But,  if   we  c.n   only  put  three   tens  into  one,  as  often   at 


lintte,  Montana,  each  ton  of  ore  must  be  charged  with  16  for 
matte-refiniug,  making  a  difference  in  results  of  $1,000  a  day  for 
a  smelter  of  ordinary  capacity. 

The  main  factor  in  determining  the  limit  of  concentration  is 
the  percentage  of  copper  contained  in  the  original  ore.  In  Butte 
it  is  found  more  profitable  (or  more  rapid)  to  submit  the  low-grade 
ores  to  a  mechanical  concentration  by  water,  so  that  the  material 
that  goes  to  the  furnace  will  already  assay  10  per  cent,  to  20  per 
cent,  copper. 

Experience  has  shown  that  we  cannot  make  a  product  at  the 
first  fusion  going  higher  than  50  per  cent,  to  60  per  cent,  copper, 
without  too  great  a  loss  of  metal  in  the  slag,  and  other  technical 
difficulties.     Hence,  the  low  ratio  of  concentration  at  Butte. 

The  opposite  extreme  may  be  illustrated  by  the  practice  at  the 
Argo  works  in  Colorado.  This  is,  commercially  speaking,  a  gold 
and  silver  smelter,  making  use  of  a  very  small  percentage  of  copper 
to  collect  the  precious  metals  into  a  rich  matte.  Regarded  metallur- 
gically,  however,  it  is  strictly  a  copper  smelter;  for  the  minute 
percentage  of  silver  and  gold  present  have  no  chemical  influence 
upon  the  operation.  Therefore,  we  may  regard  Argo  as  a  copper 
smelter,  treating  ores  averaging  3  per  cent,  copper,  and,  in  a  single 
fusion,  concentrating  12  or  more  tons  of  ore  into  one  ton  of  40  per 
cent,  matte.  Hence  the  possibility  of  tlie  long  and  intricate  series 
of  operations  by  which  the  silver,  gold,  and  copper  are  separated 
and  refined.  If  it  were  not  for  the  unusual  degree  of  concentra- 
tion at  the  first  smelting,  this  practice  would  not  be  a  commercial 
success;  and  if  it  were  not  for  the  low  tenor  of  the  charge  in  cop- 
per, the  high  concentration  would  be  impossible.  Therefore,  Argo 
is  not  a  purchaser  of  rich  copper  ores,  unless  they  are  very  high  in 
the  precious  metals  as  well. 

I  desire  to  particularly  call  attention  to  the  fact  that,  loWi  a 
proper  slag,  silver  and  gold  may  be  concentrated  in  matte  to  any 
reasonable  extent  (by  keeping  the  slag  siliceous  and  tolerably  free 
from  zinc,  I  have  gone  up  to  30  ounces  gold  and  2,500  ounces 
silver  per  ton  of  matte,  without  any  marked  loss),  as  they  do  not 
increase  the  bulk  of  the  matte  or,  practically  speaking,  the  per- 
centage of  its  metallic  contents,  and  thus  lessen  the  percentage  of 
the  protecting  sulphur  to  a  dangerous  degree;  but,  that  the  con- 
centration of  copper  is  limited  by  a  figure  represented  by  the  per- 
centage of  that  metal  in  the  highest  profitable  matte  that  we  dare 
to  make,(that  figure  usually  varying  from  35  per  cent,  to  60  per 


cent.,  divided  by  tlie  percentage  of  copper  iu  the  ore  smelted), less 
one-half  to  one  per  cent,  for  losses  in  the  slag. 

For  instance,  the  ratio  of  concentration  for  an  8  per  cent,  copper 
ore,  under  conditions  where  it  was  most  profitable  to  make  a  45 
per  cent,  matte,  would  be 


That  is  to  say,  six  tons  of  8  per  cent,  ore  must  be  smelted  to  pro- 
duce one  ton  of  45  per  cent,  matte. 

It  happens,  therefore,  not  infrequently,  that  there  are  mines  in 
remote  and  inaccessible  districts,  which  would  be  sufficiently  rich 
in  gold  and  silver  to  yield  good  profits,  were  it  not  that  they  were 
too  rich  in  copper.  The  rate  of  concentration  obtainable  by  smelt- 
ing is  too  low  to  yield  a  product  of  sufficient  value  to  pay  the  verj 
high  transportation  charges. 

The  principal  aim  of  the  copper  smelter  is  to  get  as  much  of  his 
ore  over  the  dump,  in  the  shape  of  slag  from  the  first  fusion,  and 
to  concentrate  his  copper,  gold,  and  silver  into  a  high-grade  matte, 
as  rapidly  and  perfectly  as  possible.  But  there  are  many  compli- 
cated chemical  changes  that  must  take  place  in  the  furnace  before 
this  result  is  obtained,  and  without  a  fair  knowledge  of  these  im- 
portant reactions  and  of  certain  of  the  laws  of  chemical  affinity, 
the  smelter  cannot  have  any  sound  insight  into  his  work,  nor  any 
certainty  of  succeeding  when  he  is  confronted  with  new  ores  or 
untried  conditions. 

Old  smelters,  who  pride  themselves  on  being  "practical,"  should 
realize  that  "practical  men"  usually  have  infinitely  more  theories 
on  every  subject  than  scientific  men;  only  they  are  all  wrong. 

The  most  important  reactions  that  occur  in  the  furnace  will  be 
briefly  enumerated  in  the  description  of  each  method  of  smelting. 

The  ordinary  products  of  copper  furnaces  may  be  blister  copper, 
black  copper,  copper  bottoms,  matte,  speiss,  slag,  and  flue-dust. 

There  are  various  excellent  metallurgical  works  in  which  these 
substances  are  thoroughly  discussed  and  analyzed.  I  shall,  there- 
fore, merely  offer  some  few  practical  observations  about  them  that 
do  not  find  a  place  in  the  ordinary  text-books. 

Blister  copper,  or  more  properly,  blistered  copper,  is  a  high- 
grade  crude  copper  iu  which  nearly  all  the  oxidizable  impurities 
have  been  removed  by  slagging  and  volatilization.  Good  blister 
contains  from  97  per  cent,  to  99  per  cent,  copper  and  only  0.25 
per  cent,  to  0.75  percent,  sulphur,  which,  at  the  high  melting  point 

THE    SMELTIKO^    OF    COPPER.  327 

of  metallic  copper,  aud  in  the  presence  of  air,  escapes  rapidly  as  sul- 
phurous, and  anhydrous  sulphuric  acid  gas.  This  ebullition  of 
gas  continues  up  to  the  moment  of  chilling,  and  the  gas  still  gen- 
erated in  the  molten  portion  of  the  pig  raises  little  bubbles  and 
blisters  on  the  surface  of  the  metal,  whence  its  name  is  derived. 

As  may  be  inferred,  the  production  of  this  material  is  usually 
conljued  to  operations  conducted  with  a  powerfully  oxidizing 
atmosphere,  such  as  reverberatory  furnaces  aud  Bessemer  con- 
verters. It  may,  however,  under  exceptional  conditions,  be  pro- 
duced in  blast-furnaces  ruuuing  on  oxidized  ores,  and,  as  an 
experiment,  I  have  produced  excellent  blister  from  roasted  matte, 
in  the  little  black  copper  cupolas  at  Ely,  Vermont,  which,  for  the 
past  30  years  have  been  run  something  after  the  fashion  of  a  pyritic 
smelter,  with  a  highly  oxidizing  atmosphere,  aud  producing,  ordi- 
narily, black  copper  of  the  highest  grade.  I  have  seen  excellent 
blister  copper  produced  by  Dr.  Trippel  from  oxidized  ore  in  the 
Longfellow  cupolas.  This  product,  when  broken,  has  the  true 
rosy  color  of  pure  copper,  but  not  its  fine,  silken  texture. 

It  is  very  tough  when  cold,  but  its  quality  of  redshortness  enables 
the  smelter  to  separate  the  pigs  of  a  bed  of  blister  as  tapped  from 
the  furnace,  by  breaking  the  narrow  necks  that  still  connect  the 
pigs,  the  instant  tliat  they  are  sufficiently  set  to  stand  the  pressure 
of  the  bars  used  in  prying  them  apart. 

Black  copper  is  the  name  given  to  the  more  or  less  impure 
metallic  copper  produced  in  blast-furnaces  when  running  on  oxide 
ores  or  roasted  sulphide  material.  It  is  always  an  alloy  of  copper 
with  one  or  more  other  metals,  generally  containing  several  per 
cent,  of  iron,  often  lead,  and  many  other  impurities,  according  to 
the  ores  from  which  it  is  produced.  It  usually  contains  1  per  cent. 
to  3  per  cent,  or  more  sulphur.  On  cooling,  the  surface  oxidizes, 
giving  it  a  dull,  blackish  appearance,  nor  does  its  fracture  show 
either  the  exact  color  or  texture  of  pure  copper. 

Copper  bottoms  is  a  technical  expression,  referring  to  a  metallic 
product  of  a  very  indefinite  composition,  made  (usually)  iu  rever- 
bei'atory  furnaces  by  smelting  rich  cupriferous  substances  without 
sufficient  sulphur  to  quite  satisfy  the  copper  present.  The  affinity 
of  metallic  copper  for  certain  substances  is  much  greater  than  that 
of  copper  matte,  aud  the  object  of  employing  this  smelting  for 
** bottoms"  is  to  cause  these  substances  to  combine  with  a  small 
fraction  of  metallic  copper,  by  which  the  main  portion  of  the  cop- 
per is  obtained  in  a  matte  freed  from  them.     These  alloying  sub- 


Stances  may  be  objectionable,  as  arsenic,  antimony,  tin,  lead, 
tellurium,  etc.,  or  may  be  highly  desirable,  as  gold  or  silver. 

Matte  (regulus)  is  ordinarily  the  main  valuable  product  in  the 
first  fusion  of  sulphide  ores  of  copper.  Although  every  metallur- 
gist is  extremely  familiar  with  this  curious  substance,  I  am  at  a 
loss  how  to  define  it,  as  it  has  but  a  single  essential  constituent — 
sulphur.  Without  sulphur  we  cannot  have  a  matte  in  the  sense 
in  which  this  term  is  commonly  understood.  The  copper  metal- 
lurgist would  naturally  consider  copper  a  rather  indispensable  con- 
stituent of  his  matte,  but  the  gold  and  silver  sulphide-smelter 
migiit  make  a  matte  containing  no  trace  of  copper,  or,  possibly, 
no  iron.  Nickel,  cobalt,  lead,  or  bismuth  may  take  the  place  of 
either  or  both  of  the  metals  just  mentioned;  manganese  or  zinc 
may  replace  them  to  a  marked  extent,  while  those  metallurgists 
accustomed  to  running  heavy-spar  ores  in  cupolas  need  scarcely  be 
informed  that  sulphide  of  barium  may  become  a  constituent  of  the 
matte  to  an  almost  unlimited  extent. 

But,  for  the  purposes  of  the  copper  smelter,  matte  may  be  gen- 
erally regarded  as  a  mixture  of  cuprous  sulphide  (CujS)  with  fer- 
rous monosnlphide  (FeS)  in  varying  proportions.  Thus,  in  rapid 
blast-furnace  smelting  in  a  cupola  with  boshes,  where  the  material 
is  calcined  ores,  or  ores  containing  no  bisulphides,  and  where  we 
can  pretty  nearly  disregard  any  volatilization  or  oxidation  of  the 
sulphur  in  the  furnace  itself,  we  may  consider  that  each  pound  of 
copper  present  will  take  up  one-fourth  of  a  pound  of  sulphur,  and 
that  the  remaining  sulphur  will  take  up  iron  at  the  rate  of  about 
one  and  three-fourths  pounds  for  each  pound  of  sulphur,  all  these 
newly  produced  sulphides  mixing  together  to  form  a  more  or  less 
homogeneous  matte. 

In  less  rapid  smelting,  and  where  the  volume  of  blast  is  great, 
and  the  shape  of  the  furnace  such  as  to  favor  oxidation,  the 
amount  of  sulphur  eliminated  as  sulphurous  acid  may  be  very 
great.*  But  in  steady  running,  we  can  usually  determine  pretty 
closely  our  co-efficient  of  oxidation  in  each  individual  case,  and 
should  thus  be  able  to  determine  quite  accurately  the  grade  of  our 
matte  in  advance,  were  it  not  for  the  possible  presence  of  a  dis- 

*It  is  this  fact  that  puts  into  our  hands  the  power  of  controlling  the  rate  of 
concentration  in  blast-furnace  smelting.  This  fact  has  been  long  and  loudlv 
insisted  upon  bv  F.  L.  Bartlett  and  Herbert  Lang,  but,  apart  from  the  pyritic 
smelters,  has  apparently  found  few  receptive  listeners.  It  will  be  more  fully 
discussed  in  other  chapters. 

THE    SMELTING    OF    COPPER.  229 

turbing  element  that  is  so  curious  and  unexpected  as  to  cause 
many  metallurgists  to  deny  the  possibility  of  its  existence,  until  care- 
ful and  repeated  investigations  seem  to  have  settled  the  question. 
This  unlooked  for  substance  is  magnetic  oxide  of  iron,  which  is  a 
frequent,  and  occasionally  important,  constituent  of  mattes.  It 
behaves  in  a  manner  that  appears  at  the  first  glance  somewhat 
paradoxical,  for  it  seems  to  be  formed  most  persistently  and  in  the 
greatest  quantities  in  furnaces  where  there  is  the  strongest  reduc- 
ing action,  and  where  either  a  contracted  hearth  and  considerable 
height  of  ore  column,  or  a  large  proportion  of  sulphur  in  the 
charge,  would  seem  to  forbid  the  possibility  of  any  oxidizing  influ- 
ence. I  have  frequently  found  it  in  considerable  amounts  in  the 
matte  produced  by  the  rapid  smelting  of  partly  oxidized  ores  in 
the  large  type  of  Rachette  furnaces,  and  have  noticed  it  iu  still 
greater  proportion  in  the  low-grade  matte  produced  during  the 
quick  fusion  of  siliceous,  raw  pyrites  fines,  the  charge  containing 
25  per  cent,  to  30  per  cent,  sulphur.  It  also  frequently  occurs  in 
lead-furnace  mattes  in  spite  of  the  powerful  reducing  action  result- 
ing from  slow  smelting,  high  ore-column,  and  contraction  of  the 
shaft  at  the  tuyeres. 

Certain  observations  of  W.  L.  Austin  first  assisted  me  in  ex- 
plaining this  phenomenon — to  my  own  satisfaction  at  least.  Aus- 
tin noticed  that  in  practising  pyritic  smelting  with  small  tuyere? 
and  a  high  blast  pressure,  the  partially,  or  entirely,  molten  sul- 
phides, as  they  dropped  in  front  of  the  tuyeres  and  received  the 
full  force  of  the  blast,  were  often  in  part  changed  to  magnetic 
oxide,  a  cauliflower-like  excrescence  of  this  oxide  forming 
almost  instantaneously  on  the  surface  of  a  partially  fused  mass, 
and  this  in  spite  of  the  proximity  of  a  great  preponderance  of 
vaporous  sulphur  and  sulphurous  acid.  This  may  well  be  the 
origin  of  much  of  the  magnetic  oxide  in  the  instances  that  have 
come  under  my  own  notice.  Being  a  feeble  base  and  of  high  spe- 
cific gravity,  it  does  uot  combine  with  the  silica,  but  settles  to  the 
bottom,  mixing  with  the  matte  and  becoming  a  part  of  the  latter. 
This  formation  of  magnetic  oxide  of  iron  is  generally  an  unfortu- 
nate circumstance,  doing  harm  in  at  least  flve  different  ways: 

1.  It  robs  the  slag  of  the  iron  that  is  needed  for  flux. 

2.  It  lessens  the  dissolving  power  of  the  matte  for  silver,  and 
perhaps  for  gold. 

3.  It  increases  the  quantity  of  matte  tro  be  treated  later. 


4.  It  makes  the  matte  exceedingly  tough  and  tenacious,  and 
expensive  to  break  or  pulverize. 

5.  It  makes  the  charge  less  fusible. 

If  our  theory  of  this  formation  of  magnetic  oxide  of  iron  be  cor- 
rect, it  is  very  easy  to  suggest  the  remedy.  It  is  not  too  rapid 
nor  too  slow  fusion,  nor  too  much  nor  too  little  reduction  that 
causes  the  formation  of  magnetic  oxide.  It  is  simply  too  high 
wind  pressure;  and  that  this  circumstance  seems  to  stand  in  close 
relation  to  its  production  is  shown  by  the  fact  that,  in  the  cases 
that  I  have  just  referred  to,  the  production  of  this  unwelcome 
oxide  diminished  greatly,  or  ceased  completely,  with  the  lessening 
of  the  blast  pressure.  But  this  modification  of  practice  means 
something  more  than  simply  reducing  the  blast  jjressure;  for  if 
this  alone  were  done,  the  capacitj  of  the  furnace  would  piobably 
fall  off  to  an  extent  that  could  not  be  tolerated.  The  powerful 
blast  that  was  used  conduced  to  rapid  smelting  and  great  capacity, 
and  also  presupposed  tolerably  small  tuyeres  and  a  furnace  shaft 
of  considerable  diameter,  or  width;  probably  40  to  48  inches. 
The  weakened  blast  now  proposed  cannot  successfully  penetrate 
the  ore  column  in  a  shaft  over  34  inches  in  width,  and  this  may, 
in  some  cases,  much  better  be  reduced  to  30  inches,  and  the  proper 
capacity  retained  by  enlarging  the  furnace  in  the  only  dimension 
possible,  that  of  its  length. 

This  gives  us  a  loug,  narrow  rectangle,  and,  as  we  are  obliged 
to  decrease  our  wind  pressure,  we  must  enlarge  our  tuyeres,  in 
order  to  obtain  a  sufficient  volume  of  air  to  burn  the  considerable 
quantities  of  fuel  that  fill  this  space.  The  low  pressure  and  large 
volume  of  blast  required  suggest  at  once  the  employment  of  a 
large  fan  blower  in  place  of  a  positive,  or  semi-positive,  blast 
machine,  and,  if  it  were  not  for  the  annoyance  caused  by  large 
belts  driving  small  pulleys  at  a  high  speed,  I  should  feel  much 
inclined  to  return  to  the  stand  taken  some  years  ago  by  Mr.  H.  M. 
Howe  in  regard  to  fan-blowers. 

In  a  work  like  the  present  one,  devoted  almost  exclusivelv  to 
the  practical  side  of  metallurgy,  it  is  impossible  to  even  enumerate 
all  the  interesting  questions  still  presented  by  matte  for  study  and 

Is  it  a  chemical  combination,  a  mixture,  or  a  partial  alloy? 

What  are  the  affinities  of  the  various  sulphides  that  it  may  con- 
tain, at  smelting  temperatures,  and  how  do  they  vary  among 
themselves  as  the  temperature  rises  and  sinks? 


What  affinity  or  power  of  alliage  is  there  between  the  metallic 
sulphides  and  those  of  barinni  and  calcium? 

Why  does  the  same  matte  separate  more  quickly  and  thoroughly 
from  an  acid  slag  than  from  an  equally  light,  and  much  thinner, 
basic  slag  (containing  principally  alkaline  and  earthy  bases)? 

Why  does  the  capacity  of  matte  to  collect  the  silver  of  an  ordi- 
nary charge  increase  to  a  certain  point  as  its  copper  contents 
increase,  and  then  retrograde  as  the  matte  becomes  still  richer  in 
copper,  while  its  affinity  for  gold  continues  increasing,  metallic 
copper  having  the  greatest  affinity  of  all? 

Why  does  a  cone  of  matte,  allowed  to  cool  naturally,  crack  par- 
allel with  its  surface  when  containing  over  50  per  cent,  copper, 
and  at  right  angles  to  this  direction  when  below  50  per  cent. 

These  are  but  a  few  of  the  unexplained  phenomena  regarding 
matte  that  are  constantly  forcing  themselves  on  the  copper  smelter's 

Speiss,  as  ordinarily  understood,  is  a  basic  arsenide,  or  antimo- 
nide  of  iron,  often  with  nickel,  cobalt,  lead,  bismuth,  copper,  etc., 
having  a  metallic  luster,  high  specific  gravity,  and  a  strong  ten- 
dency toward  crystallization.  It  takes  up  gold  with  avidity,  but 
has  a  less  affinity  for  silver  than  copper  matte  has. 

It  hr.::  always  seemed  to  me  that  here  is  a  field  that  has  not  been 
sufficiently  exploited.  Especially  since  bessemerizing  and  pyritic 
smelting  are  becoming  so  important,  it  is  worth  while  to  consider 
to  what  degr'^^,  and  with  what  advantages,  speiss  may  be  used  to 
replace  sulphides  under  favorable  conditions.  We  have  several 
instances  where  it  has  been  used  to  collect  silver,  gold,  or  copper. 
A  late  notable  example  in  the  Transvaal,  South  Africa,  of  which, 
I  regret  to  say,  I  have  no  personal  knowledge,  is  described  by  Mr. 
W.  Bettel  in  the  Chemical  News  of  June  2(i,  1891.  He  describes 
the  production  of  an  argentiferous,  antimonial  coiiper  speiss  of  the 
following  composition,  from  smelting  oxidized,  ferruginous  oros. 
containing  much  antimonate  of  iron,  and  4  per  cent,  of  copper  in 
the  shape  of  carbonates,  and  36  ounces  silver  per  ton  (0.123  per 

*  This  fact  was  first  pointed  out  to  me  by  H.  C.  Bellinger  at  the  Montana  Ore 
Purchasing  Company's  smelter  at  Butte,  Montana. 


Copper 52.50 

Antimony 3S.0O 

Arsenic 2.00 

Sulphur 2.06 

Iron 3.60 

Silver 159 

Lead 0.25 


The  ore  is  smelted  in  reverberatory  furnaces,  and  some  91  per 
cent,  of  the  silver  and  copper  is  collected  in  the  speiss.  The  con- 
centration averages  16.4  tons  into  one. 

Slags. — The  copper  metallurgist  approaches  this  subject  from  a 
totally  different  standpoint  from  that  of  the  lead-silver  smelter. 
It  has  been  shown  by  many  able  writers  that  to  oljtain  slags  low  in 
lead  and  silver,  it  is  advisable  in  lead  smelting  to  form  the  slag  so 
that  there  may  be  some  definite  and  constant  ratio  between  the 
iron,  lime,  and  silica  that  form  its  principal  constituents.  After 
numerous  experiments  under  varying  conditions,  I  am  unable  to 
detect  any  such  law  that  can  be  applied  to  copper  n^atte  slags. 
From  a  considerable  number  of  determinations,  I  select  the  fol- 
lowing, the  chemical  vpork  of  these  experiments  having  been 
mostly  done  by  Messrs.  D.  Murphy,  A.  R.  Vincent,  and  T.  G. 
Rockwell.  In  all  the  cases  the  sampling  was  conducted  with  care, 
a  small  ladlefnl  of  slag  being  caught  under  the  slag-spout  just  as 
each  pot  was  pulled  away,  while  equal  pains  were  taken  to  obtain 
a  true  sample  of  the  matte.  Each  separate  type  of  slag  was  run 
for  six  hours,  and  no  samples  were  taken  of  molten  material  from 
the  fresh  charge  until  it  had  been  in  the  furnace  double  the  time 
necessary  to  reach  the  tuyeres.  Then  the  furnace  and  forehearth 
were  tapped  completely  dry,  and  sampling  was  begun  after  the 
fresh  flow  of  products  had  become  well  established.  The  furnace 
part  of  such  experiments  is  very  easily  and  cheaply  done,  as  it  is 
oulv  necessary  to  add  or  subtract  a  certain  calculated  portion  of 
siliceous,  or  basic  ore,  at  each  charge. 







FeO.  |BaO;Cu.^SiCu. 

I  Normal  charge.. 
^Increased  silica.. 
Increased  silica.. 
iDirainished  silica 
I  Normal  charge.. 
Increased  silica  . 
Increased  silica., 
ilncreased  silica., 
i Increased  silica.. 
Diminished  silica 

Siliceous  roasted  ore  and  roasted 
-concentrates ■ 




2  .57.6 
6    64.4 

1  63.1 

3  61.1 
6  57.7 
3    55.5 

2  undet 

I   7.2  0.;*i3.5   m 

6.1,0.31  3.15  39, 
'   8.6,0.44  2.9   4:i, 

12.2  0.78  4.4  ,35. 
CaO0.61  0.71  54. 

....  0.820.91  56. 

....0.770.8    56. 

....  0.61  0.56  58, 

....0.6    0.48  59. 

....  0.98  1.86,50. 



9  601 
6  666 




4l  31. 
5    22. 

Rich,  siliceous,  argentiferous  and  I 
zinciferous  dry  ores,  roasted  ar-J 
gentiferous  pyrite,  and  a  lime- 1 
stone  containing  copper  glance. .  ( 

! Normal  charge..!  .32. 5 1  .52.1 
■Increased  silica..  35. 4|  .50.7 
Increased  silica..!  39.3,  46.9 

11.7  0.31  1.9  !32.7 
9.4  0.3  2.2  2S.6  125.5 

1  3.57 




4  .... 
4  .... 




In  the  above  table  every thiug  is  given  in  percentages,  excepting 
the  gold  and  silver.  These  are  given  in  ounces  per  ton  of  2,000 
pounds.  To  reduce  this  to  percentage,  multiply  the  ounces  per 
ton  by  0.003436. 

The  above  results  were  selected  for  publication  as  being  among 
the  most  uniform  and  complete  of  a  considerable  series  of  similar 
tests,  but  I  can  detect  nothing  in  them,  or  in  any  of  the  figures 
obtained,  to  show  that  the  freedom  of  a  copper  slag  from  valuable 
metals  stands  in  any  especial  relation  to  the  stochiometrical  pro- 
portion or  arrangement  of  its  constituents. 

We  feel,  therefore,  comparatively  untrammeled  as  to  the  compo- 
sition of  our  slags,  providing  always  that  they  are  sufficiently  fusi- 
ble and  that  their  specific  gravity  is  not  so  great  as  to  hinder  the 
settling  out  of  them  of  the  matte  particles.  In  planning  a  new 
slag,  we  are,  within  reasonable  limits,  guided  by  commercial  rather 
than  by  chemical  influences,  and  are  tolerably  independent  of  the 
limestone  quarry.  That  this  is  peculiarly  the  case  in  Pyritic 
Smelting  will  be  seen  when  that  subject  is  reached.  Nearly  every 
copper  metallurgist  begins  his  furnace  work  by  trying  to  make  as 
basic  and  ferruginous  a  slag  as  circumstances  will  permit,  and  fin- 
ishes by  making  his  slags  as  siliceous  as  possible.  While  skill  and 
good  settling  facilities  may  succeed  in  making  a  tolerably  clean 
slag  from  a  basic  charge,  it  is  very  ranch  easier  and  surer,  and 
need  not  necessarily  take  a  pound  more  of  coke,  to  make  a  quite 
siliceous  slag.  This  is  especially  the  case  where  copper  is  scarce, 
and  the  minute  proportions  of  tellurium,  bismuth,  and  other  com- 


paratively  nnstudied  substances  that  so  increase  the  power  of  the 
matte  to  collect  the  precious  metals,  are  wanting.  So  far  as  ni} 
own  experience  goes,  I  consider  an  acid  slag  in  such  a  case,  an 
absolute  si?ic  gnu  non. 

In  the  smelting  of  sulphide  ores,  unless  some  unusual  conditions 
prevail,  the  copper,  silver,  and  gold  contained  in  the  slag  are  pres- 
ent in  the  shape  of  shots  or  prills  of  matte.  Most  of  these  particles 
are  extremely  minute  and  can  be  best  seen  by  reflected  light,  and 
with  the  aid  of  a  good  magnifier.  There  is  no  excuse  for  this  con- 
dition of  things,  if  it  at  all  exceeds  the  customary  limits.  Either 
the  slag  must  be  unsuitable  in  consistency  or  gravity  for  the  sepa- 
ration of  the  matte  globules,  or,  what  is  very  much  more  common, 
the  settling  facilities  are  inadequate.  Especial  attention  will  be 
paid  to  this  important  subject  when  we  come  to  consider  the  con- 
struction of  furnaces. 

"  What  is  the  best  slag  to  make  under  my  conditions?"  is  rather 
a  commercial  than  a  metallurgical  question.  Pretty  much  any- 
thing, within  wide  limits,  can  be  smelted,  and  if  it  is  more  profit- 
able to  produce  a  slag  containing  2  per  cent,  copper  and  10  ounces 
silver  than  it  is  to  flux  the  charge  so  as  to  save  those  metals,  the 
former  is  the  proper  slag  to  make.  These  abnormal  conditions 
become  more  and  more  rare  as  the  Western  country  is  opened  up 
by  railroads,  but  they  still  exist;  and  in  portions  of  Mexico  may 
continue  to  prevail  for  many  years. 

The  usual  object  of  smelting  a  copper  ore  is  simply  to  divide  it 
into  two  portions:  a  small  quantity  of  matte  for  further  treatment, 
and  a  large  amount  of  slag  to  go  over  the  dump.  Now  it  is  entirely 
immaterial  how  this  object  is  accomplished,  or  whether  the  ore 
has  been  thoroughly  fused  or  only  half  melted,  providing  that  the 
work  has  been  done  in  the  cheapest,  quickest,  and  most  effective 
way  possible  under  the  circumstances.  For  instance,  the  Swansea 
smelters  long  ago  found  out  tliat  it  did  not  pay  them  to  flux  all 
the  silica  when  running  on  a  highly  qnartzose  charge.  A  rever- 
beratory  slag  may  contain  close  on  to  50  per  cent,  of  nnmelted  frag- 
ments of  pure  quartz,  and  yet  be  clean  and  satisfactory;  the  main 
requirement  being  that  there  shall  be  a  sufficient  proportion  of 
molten  slag  to  float  the  un fused  particles,  and  enable  the  worthless 
portion  of  the  charge  to  be  dragged  out  of  the  furnace  without 
carrying  with  it  the  valuable  part.  This  species  of  liouation  may 
at  times  be  used  to  great  advantage. 

Flue-dust. — The  main  practical  interest  attached  to  this  prod oct 


is  connected  with  the  methods  for  its  collection  and  treatment, 
which  are  considered  elsewhere. 

For  practical  purposes  we  may  distinguish  three  totally  separate 
and  distinct  methods  of  smelting: 

(a)  Blast-furnace  smelting  with  carbonaceous  fuel.  Suited  to 
every  class  of  copper  ore,  whether  metallic,  oxides,  or  sulphides. 
Atmosphere  in  furnace,  reducin'g. 

{b)  Reverberatory  smelting.  Mainly  for  sulphides.  In  a  sub- 
ordinate degree,  for  metallic,  and  oxide  ores.  Atmosphere  in 
furnace,  neutral. 

(c)  Pyritic  smelting.*  For  sulphide  ores,  though  oxide,  or 
metallic  ores  may  always  be  added  when  there  is  an  excess  of  sul- 
phide.    Atmosphere  of  furnace,  oxidizing. 

*  By  the  terra  "  Pyritic  Smelting,"  I  intend  to  designate  that  distinct  and 
characteristic  process  by  which  sulphide  ores  are  smelted,  in  the  main,  without 
the  use  of  carbonaceous  fuel,  the  necessary  heat  for  the  operation  of  smelting 
being  obtained  from  the  combustion  of  the  sulphur  and  iron  contained  in  the 
ore  itself.     See  chapters  xiv  and  xv. 



The  one  distinctive  feature  of  the  blast-farnace  is  tlie  absence 
of  a  separate  fireplace,  the  ore  and  fuel  being  in  direct  contact  in 
their  passage  through  the  furnace.  It  is  also,  in  a  more  general 
way,  characteristic  of  it,  that  its  operation  is  continuous,  and  that 
it  is  provided  with  a  forced  blast. 

This  rapid  combustion  of  carbonaceous  fuel  produces  a  strongly 
reducing  atmosphere,  and  brings  about  a  series  of  reactions  that, 
although  possessing  much  similarity  to  those  that  occur  in  the 
neutral  atmosphere  of  the  reverberatory  furnace  and  in  the  oxidiz- 
ing atmosphere  of  the  pyritic  smelter,  yet  differ  in  some  few  details 
that  are  of  the  utmost  commercial  importance  to  the  smelter  of 
copper  ores. 

A  thorough  familiarity  with  the  chemical  reactions  that  occur 
in  the  operations  of  calcining  and  smelting  is,  next  to  natural 
common  sense,  the  most  important  attribute  of  the  metallurgist.* 

In  describing  the  most  striking  reactions  that  take  place  in  the 
blast  furnace,  we  may  assume  the  charge  to  consist  of  calcined 
pyritous  ores  of  copper,  containing  a  little  gold  and  silver,  and 
sufficient  iron,  lime,  and  silica  to  make  a  proper  slag.  The  fuel 
shall  be  ordinary  coke,  though,  in  the  main,  the  same  reactions 
will  occur  with  charcoal. 

*  In  attempting  to  make  the  most  important  of  these  reactions  clear  to  those 
who  have  not  had  a  scientific  training,  I  must  necessarily  speak  in  general 
terms,  avoiding  such  considerations  as  the  influence  exerted  by  the  ash  of 
the  fuel,  the  frequent  occurrence  of  oxides  of  iron  in  the  matte,  the  presence  of 
sulphides  of  calcium  and  iron  in  the  slag,  the  imperfect  working  of  Fournet's 
law  regarding  the  order  of  affinity  for  sulphur  possessed  by  the  various 
metals,  etc.     These  matters  will  be  treated  of  in  their  appropriate  place. 


We  may  divide  the  constituents  of  the  ore  into  four  classes: 

Bases.  Protecting  Agents.  Reducing  Agents.  Acids. 

Iron.  Sulphur.  Coke.  Silica. 

Lime.  (As.)  (Sulphur.)  (A1,0,,V 

Copper.  (Sb.) 

(MnO.)  (Te.) 







The  tendency  of  the  reactions  that  occur  is  for  the  bases  eitlier 
to  be  reduced  to  a  metallic  condition  and  to  be  separated  out  as 
metals,  or  to  become  oxidized  and  to  combine  with  the  silica  to 
form  a  slag. 

Neither  of  these  conditions  would  satisfy  the  smelter;  for  in  the 
one  case  he  would  obtain  an  alloy  of  metallic  iron  and  copper 
(lime  being  a  very  powerful  base,  in  the  main  runs  no  risk  of  being 
reduced,  but  combines  with  silica  without  any  particular  care  being 
required),  while  the  other  alternative  would  be  that  he  would 
oxidize  and  slag  a  considerable  portion  of  the  copjier  present.  It 
is  essential,  therefore,  to  steer  a  middle  course  between  construct- 
ing and  running  a  furnace  in  such  a  fashion  that  its  reducing 
action  shall  be  powerful  enough  to  reduce  the  iron,  as  well  as  the 
desired  copper,  to  a  metallic  condition;  and  the  opposite  extreme 
(theoretical),  where  much  of  the  copper  would  be  oxidized  and 
slagged.  To  maintain  this  delicate  equilibrium  would  be  some- 
what difficult  (though  tolerably  attained  in  the  smelting  of  purely 
oxidized  ores),  were  it  not  for  the  presence  of  a  powerful  regulating 
and  protecting  agent,  in  the  shape  of  sulphur. 

This  element  has  a  very  strong  affinity  for  copper,  but,  under 
the  circumstances  that  we  are  considering,  can  only  combine  with 
it  when  the  copper  is  in  a  metallic  condition.  Therefore,  the  sul- 
phur that  is  still  present  in  considerable  quantity  in  the  imper- 
fectly calcined  ore,  aided  by  the  powerful  reducing  gases  resulting 
from  the  burning  of  the  coke,  reduces  to  its  metallic  condition 
such  copper  as  is  present  in  any  oxidized  form.  In  so  doing,  a 
portion  of  the  sulphur  itself  is  burned  by  the  oxygen  that  it  takes 
from  the  copper,  and  escapes  as  a  gas  (SOg).  The  carbonic  oxide 
is  the  main  reducing  agent,  but  it  is  very  desirable  to  be  famili:ii' 


also  with  the  reducing  elfect  of  sulphur  upou  metallic  oxides,  from 
which  springs  the  important  metallurgical  principle,  that  sulphur 
and  oxide  of  copper  smelted  together  yield  metallic  copper  and 
sulphurous  acid  gas. 

Tlie  remainder  of  the  sulphur  combines  with  the  copper  that 
has  just  been  reduced  to  a  metal,  taking  it  up  in  about  the  pro- 
portion of  four  pounds  of  copper  to  one  pound  of  snlphnr.  If 
only  enough  sulphur  were  present  to  exactly  satisfy  the  copper, 
the  resulting  matte  would  be  a  pure  subsulphide  of  copper  (OujS), 
containing  80  per  cent,  and  20  per  cent,  sulphur.  The  produc- 
tion of  so  high  grade  a  matte  would  not  only  make  the  slag  too 
rich  in  copper,  but  would  render  the  management  of  the  furnace 
more  difficult;  for  the  constant  presence  of  a  considerable  quantity 
of  matte  of  a  moderate  tenor  in  copper  keeps  the  furnace  and  fore- 
hearth  open  and  hot,  and  facilitates  rapid  driving. 

In  ordinary  work,  there  is  no  danger  of  any  such  contingency 
as  the  calcination  of  sulphide  ores  is  almost  invariably  under,  rather 
than  overdone. 

Hence,  there  is  nearly  always  more  sulphur  present  than  is 
needed  to  saturate  the  copper  in  the  proportion  of  one  pound  of 
sulphur  to  four  pounds  of  copper.  This  excess  of  sulphur  pro- 
ceeds to  attack  the  metal  for  which  it  has  the  next  greatest  affinity 
after  copper.  This  metal  is  iron,  which  combines  with  sulphur 
in  the  proportion  of  one  and  three-fourths  pounds  of  iron  to  one 
pound  of  sulphur.  The  resulting  monosulphide  of  iron  has  the 
property  of  mixing  with  subsulphide  of  copper  in  all  proportions; 
and  the  resulting  mixed  sulphides,  being  much  heavier  than  the 
slag,  separate  therefrom  and  sink  to  the  bottom. 

It  must  be  self-evident  that  the  grade  of  the  matte  will  depend 
upon  the  amount  of  sulphur  present;  for  after  a  certain  portion  of 
the  latter  has  been  burned  in  reducing  the  oxide  of  copper  to 
metal,  and  a  still  further  portion  has  combined  with  the  copper  to 
form  a  subsulphide,  every  pound  of  sulphur  that  is  left,  and  that 
is  not  burned  in  some  way,  will  take  up  one  and  three-fourths 
pounds  of  iron;  thus  diluting  the  matte  to  the  extent  of  two  and 
three-quarters  pounds  of  worthless  sulphide  of  iron  for  each  pound 
of  superfluous  sulphur  present. 

We  have  already  seen  that  it  was  necessary  to  dilute  our  matte 
to  a  certain  limited  extent  with  sulphide  of  iron,  that  it  might 
not  be  too  rich.  But  any  sulphide  of  iron  in  excess  of  the  amount 
required  to  lower  the  matte  to  the  grade  that  is  found  most  ad- 


vautageons  for  our  own  local  conditious,  will  usually  mean  a  heavy 
loss  in  two  directions. 

1.  It  makes  an  excessive  quantity  of  low-grade  matte,  thus  en- 
tailing heavy  expenses  for  its  future  treatment. 

3.  It  robs  the  slag  of  the  iron  that  is  usually  needed  as  a  flux 
for  the  silica  present,  and  carries  it  into  the  matte,  where  it  is  not 

All  this  trouble  arises  from  an  excess  of  sulphur  in  the  blast- 
furnace charge,  which,  of  course,  means  that  there  has  been  an 
insufficient  calcination.  When  smelting  with  carbonaceous  fuel, 
the  secret  of  the  economical  treatment  of  sulphide  ores  lies  in  the 
calcining  furnace.* 

Thus  far  it  has  been  convenient  to  regard  matte  merely  as  a 
mixture  of  subsulphide  of  copper  with  monosulphide  of  iron.  But, 
in  practice,  we  rarely  find  its  composition  so  simple.  Indeed,  I  can- 
not give  a  definition  of  matte  that  is  at  all  satisfactory  to  myself. 
The  subsulphide  of  copper  seems  to  be  the  most  regular  and  con- 
stant basis  to  start  from,  but  this  may  be  replaced,  in  whole  or  in 
part,  by  monosulphide  of  iron,  or  by  the  sulphides  of  nickel,  cobalt, 
lead,  manganese,  or  bismuth,  while  silver,  gold,  tin,  platinum, 
iridium,  molybdenum,  and  cadmium  are  collected  in  this  substance, 
when  they  occur  in  the  ores. 

Nor  is  this  variability  confined  to  the  electro-positive  elements. 
Sulphur  is  frequently  accompanied,  or  partly  replaced,  by  arsenic, 
antimony,  tellurium,  or  selenium,  all  of  which  combine  with  the 
copper,  iron,  etc.,  forming  frequently  a  matte  of  such  complexity 
that  it  is  impossible  to  construct  any  formula  for  it,  even  after 
the  most  careful  analysis. 

Metallic  copper,  iron,  and  lead,  and  magnetic  oxide  of  iron  are 
also  found  in  mattes,  but  I  cannot  regard  them  as  proper  constitu- 
ents of  the  same.  They  seem  to  me  either  as  substances  produced 
by  certain  reactions  inside  the  furnace,  and  merely  mechanically 
mixed  with  the  matte,  or  else  to  have  been  in  combination  with 
the  sulphur,  or  other  metalloids,  during  the  time  of  fusion,  and  to 
have  separated  out  on  cooling. 

The  sulphides  of  calcium  and  barium  are  also,  according  to  my 

*  Certain  conditions  may  render  it  more  economical  to  smelt  the  ores  raw, 
and  throw  the  bulk  of  the  work  onto  the  subsequent  converter  process.  This 
may  be  regarded  as  simply  deferring  the  calcination  to  a  later  stage  of  the 
process.  It  will  be  remembered  that  we  are  not  considering  "  Pvritic  Smelt- 
ing "  at  this  time. 


observation,  merely  admixtnres,  as  they  will,  uuder  proper  condi- 
tions, separate  and  float  on  the  snrface  of  the  heavier  snlphides. 
They  do  harm  in  three  ways: 

1.  By  lessening  the  power  of  the  matte  to  dissolve  the  precious 

2.  By  lessening  the  specific  gravity  of  the  matte,  so  that  it  wiil 
not  separate  so  perfectly  from  the  slag. 

3.  By  carrying  into  the  matte,  where  they  are  not  wanted,  bases 
that  are  usually  much  needed  in  the  slag. 

Assuming  the  slag  to  be  well  melted  and  sufficiently  fluid,  the 
action  of  specific  gravity  is  the  sole  agent  which  causes  the  separa- 
tion of  the  matte  therefrom. 

Hence,  it  is  obvious  that,  other  things  being  equal,  a  heavy 
matte  and  a  light  slag  would  cause  the  least  losses  of  metal.  But 
as  we  usually  have  to  put  up  with  ferrous  oxide  as  our  principal 
base,  we  necessarily  produce  a  slag  of  too  high  a  specific  gravity 
for  the  most  favorable  separation  of  the  matte,  and  consequently 
are  obliged  to  adopt  extensive  settling  apparatus,  and  also  to  put 
up  with  a  more  or  less  serious  loss  of  values.  Yet,  as  will  be  ex- 
plained more  fully  in  its  proper  place,  much  of  this  loss  may  be 
avoided,  even  with  heavy  slags  and  a*  light  matte,  providing  that 
the  slag  is  kept  very  hot  and  liquid  during  the  settling  operation, 
that  the  particles  of  matte  are,  as  far  as  possible,  brought  in  con- 
tact with  a  larger  body  of  molten  matte  already  settled,  and  that 
sufficient  time  is  given  for  the  slow  subsidence  of  such  globules  of 
matte  as  have  escaped  the  contact  already  referred  to. 


As  I  have  received  a  considerable  number  of  requests  to  give  a 
detailed  example  of  a  convenient  method  of  calculating  a  smelting 
mixture,  [  introduce  it  at  this  point  to  illustrate  the  principles 
that  we  have  been  considering. 

It  is  a  late  actual  case,  with  figures  evened  and  simplified  a 
little,  and  although  it  refers  to  a  raw  smelting  of  the  sulphide  ores 
and  a  subsequent  oxidizing  fusion  of  the  matte,  to  fit  it  for  the 
converter  process,  it  is  peculiarly  suited  to  illustrate  the  reactions 
mentioned  in  this  chapter,  the  ore  being  unusually  simple  and 

The  ore  that  we  will  take  as  a  practical  example  shall  consist  of 
a  mixture  of  copper,  and  iron  pyrites,  in  a  slaty  and  abundant 
gangue.     A  considerable  portion  of  the  chalcopyrite  is  sufficiently 


massive  to  be  cheaply  picked  out  by  hand  as  a  siliceous  first-class 
ore.  The  remainder,  and  by  far  the  greater  proportion  of  the  ore, 
is  to  be  subjected  to  a  mechanical  concentration.  Although  it  is 
not  a  good  ore  for  the  purpose,  and  test  runs  have  shown  that  it 
will  undergo  a  heavy  loss,  its  abundance  and  extreme  cheapness  of 
mining,  and  the  lack  of  suitable  basic  ores  for  liux,  render  it 
cheaper  to  waste  a  certain  proportion  of  the  metal  than  to  save  it. 
The  occurrence  of  a  moderate  amount  of  gold  and  silver  in  the  ore 
also  bars  the  employment  of  a  wet  method.  It  would  be  superflu- 
ous to  go  into  more  detailed  explanations  of  the  reasons  for  adopt- 
ing the  method  of  treatment  to  be  discussed,  as  this  is  in  no  wise 
the  object  of  the  example. 

To  make   matters  plain  from   the  outset,  I  will  begin  with  the 
ore  as  it  is  delivered  at  the  mouth  of  the  shaft. 


We  will  assume  that  the  mine  delivers   daily  to  the  concentrator,  600  tons 
(1,200,000  pounds)  crude  ore,  averaging  4.6  per  cent,  copper. 

Pounds  Cii. 

600  tons  ore  contain , 55,200 

Products  of  hand  picking: 

Pounds  Cu. 

90  tons  waste  rock  1.5  per  cent =    2,700    (4.9  per  cent,  of  original  Cu.) 

80  tons  siliceous  selected  ore,  10  per 

cent =  16.000 

430   tons   ore  for  concentrator,  4.244 

per  cent =  36,500 

600  tons.  Total 55,200 

Products  of  concentrating  mill: 
We  start  with  430  tons  ore,4.2442  per 

cent.  Cu ; =36,500 

Loss  in  concentration  =  40  per  cent,  on  this  ore. 
We  produce 

320.5  tons  tailings,  2. 2777  per  cent.  Cu  =  14,600  (26.5  per  cent,  of  original  Cu) 
109.5  tons  concentrates,  10  per  cent.  =  21,900 

430     tons.  Total 36,500 

Total  loss  of  copper  thus  far  =  31.5  per  cent,  on  original  amount. 
The  products  to  go  to  the  smelter  are,  therefore: 

109.5  tonspyritous  concentrates,  10  per  cent.,  containing  21,900  pounds  copper 
80     tons  siliceous  selected  ore,  10  per  cent.,  containing  16,000       "  " 

189.5  tons  in  total,  containing  37,900       " 




The  smelting  treatment  is  to  consist  of  tiiree  operations: 

1.  Smelting  the  raw  (and  tolerably  granular)  concentrates,  with 
the  major  portion  of  the  raw,  siliceous  selected  ore,  and  with  tlie 
converter  and  matte-concentration  slags,  in  blast-furnaces,  witli 
coke,  for  a  matte  with  about  30  per  cent,  copper,  and  a  slag  with 
about  33  per  cent,  silica. 

2.  An  oxidizing  (pvritic)  smelting  of  the  raw  matte  produced 
in  No  1.  operation,  with  the  remainder  of  the  siliceous  selected 
ore,  in  a  blast-furnace,  to  form  a  55  per  cent,  matte  for  the  con- 
verters, and  a  moderately  basic  slag,  to  go  back  to  the  ore  cupolas. 

3.  Bessemering  the  matte  from  iNlo.  2  up  to  high  blistered  copper 
in  converters,  the  ferruginous  slag  therefrom  resulting  going  back 
to  the  ore  cupolas. 

The  tirst  calculation  required  is  to  find  out  the  exact  amounts 
of  the  substances  that  we  have  to  smelt.  The  proportion  of  pyrite, 
chalcopyrite,  and  gangue  contained  in  the  siliceous  ore  and  in  the 
concentrates  has  already  been  carefully  established.  The  two 
pyritic  minerals  are  known  to  be  practically  pure,  and  the  average 
analysis  of  the  slaty  gangue  has  been  settled. 


80  tons  (160,000  pounds)  10  per  cent.  Cu. 

t  Copper 34  per  cent. 

Copper  pyrites 29.4    per  cent,  <  Iron 31 

(sulphur 35 

jlron 47 

/Sulphur 53 

(Silica 80 

(Earths 20 


(Deduced  from  above  table.) 


,  ^Jn  29.4  per  cent,  copper  pyrites  =    9.12  per  cent. 

Iron  pyrites 22.6 

Gangue  rock 48.0 

10.0  per  cent. 

/In  22.6 

Q  ,  1         (In  29.4 

Sulphur.  <^    ^^  ^ 

^  In  22.6 

iron  pyrites 

copper  pyrites 
iron  pyrites 

=  10.60 

=  10.29 
=  11.98 



Silica  in  48  per  cent,  gangue  rock  at  80  per  cent 38.4 

Earthsin48        "  "         "  20        "      , 9.6 




109.5  tons  (219,000  pounds)  10  per  cent.  Cu. 

/  Copper 34  per  cent. 

Copper  pyrites 29.4    per  cent.  \  Iron 31 

Iron  pyrites 60.6 

■\  iruu Oi 

(Sulphur. ...  35 

jlron 47 

(Sulpliur....  53 

Gangue  rock     10.0         "  (Silica 70         " 

1  Earths 30         " 


(There  being  some  hornblende  in  the  ore,  the  gangue  is  more  basic  in  the 
concentrates  than  in  the  unwashed  ore.) 


(Deduced  from  above  table.) 
Copper 10.0  per  cent. 

Iron  i  ^"  ^^■'^  P^^  ^®'^*'  ^°PP®^  pyrites  =    9.12  per  cent. 
(In  60.6         "        iron  pyrites       =  28.48 

.  37.6 

Sulphur  . .  -1  ^"  ^^"^        "        ^^PP'^''  P>^"^^' 
(In  60.6        "        iron  pyrites 

iron  pyrites       =32.11 


Silica  in  10  per  cent,  gangue  rock  at  70  per  cent. 7.0      " 

Earths  in  10       "  "        "     at  30        "     3.0      " 

100.0      " 

Of  course  these  deductive  analyses  have  been  checked  by  many 
actual  analyses  of  large  lots  of  ores  and  concentrates  from  various 
portions  of  the  veins. 

Ore  Cupolas. — The  siliceous  ore  and  concentrates  just  described 
have  now  to  be  melted  raw  in  blast  furnaces,  with  coke,  for  a  30  per 
cent,  matte,  using  the  slags  from  the  mattte-concentration  cupola 
and  from  the  converters  as  flux.  I  will  not  attempt  to  give  the 
reasons  for  adopting  this  somewhat  peculiar  method,  whereby  there 
is  to  be  a  concentration  of  only  three  into  one  in  the  first  srueltins. 
Yet  they  are  very  simple,  when  it  is  understood  that  coke  and 
labor  are  excessively  cheap,  basic  flux  scarce,  and  that  strong 
reasons  exist  for  avoiding  a  calcining  plant. 

By  using  large  furnaces,  a  great  volume  of  blast,  and  slow  run- 
ning, there  will  be  no  difficulty  in  producing  a  30  per  cent,  (or 
much  higher)  matte  at  the  first  smelting,  and  the  heat  produced 
hv  the  combustion  of  the  raw  pyrites  in  the  furnace  will  doubtless 



bring  the  coke  consumption  somewhat  below  the  estimated  amount, 
10  per  cent. 

Owing  to  the  complications  introduced  by  smelting  a  portion  of 
the  siliceous  selected  ore  with  the  matte  in  the  second  operation, 
and  also  by  returning  the  ferruginous  slags  from  that  operation, 
and  from  the  converters,  to  the  ore  cupola,  we  cannot  calculate 
the  ore-mixture  as  a  straightforward  proposition,  but  must  begin 
by  making  some  reasonable  assumption,  in  order  to  get  at  the 
amount  of  slags  that  we  shall  have  to  resmelt.  The  slag  from  the 
ore  cupolas  should  not  carry  over  0.4  per  cent,  copper  at  the  out- 
side, and,  with  the  style  of  settlers  provided,  will  not  make  over 
one-fourth  of  one  per  cent,  of  foul  slag.  This  is  so  small  an 
amount  to  be  resraelted  that  we  may  neglect  it  entirely  in  our  cal- 
culation; nor  need  we  take  into  account  the  copper  in  the  slags 
that  are  resmelted  from  the  two  last  operations,  as  it  is  a  constant 
amount,  and  is  eventually  recovered. 

We  will  start  our  calculation  for  the  ore  cupolas  with  the  fol- 
lowing mixture: 






1  Sulphur. 

Siliceous  selected  ore 

120,000  with 
219.000  with 







339,000  with      aSLftm 




!    119,616 

To  produce  a  normal  30  per  cent,  matte  with  the  above  quantity 
of  copper,  will  use  up  sulphur  and  iron  as  follows:  33,900  pounds 
copper  will  makt*  42,375  pounds  subsulphide  of  copper  (80  per  cent, 
matte),  or  33,900  pounds  copper  will  make  113,000  pounds  of  30 
per  cent,  matte,  containing  subsulphide  of  copper,  42,375  pounds; 
sulphide  of  iron,  70,625  pounds;  a  total  of  113,000  pounds. 

I'he  iron  which  is  thus  tak^ii  up  into  the  matte  in  the  shape  of 
sulphide  of  iron  amounts  to  44,017  pounds.  Deducting  this  iron 
that  we  have  thus  temporarily  lost  from  the  total  amount  of  iron 
contained  in  tiie  mixture,  we  have  105,084  minus  44,917  equal  to 
01,067  pounds  of  iron  still  available  to  fl'ix  the  silica  of  the  mix- 
ture. This  iron  when  oxidized  to  ferrous  oxide»  so  that  it  can 
enter  the  slag,  will  wpigh  78,471  pounds. 

We  also  have  a  considerable  amount  of  earths  available  to  flnx 
the  silica,  and  as  they  consist  almost  exclusively  of  magnesia  and 
a  little  alumina,  we  may  call  them  worth  twice  as  much,  pound 
for  pound,  as  the  ferrous  oxide,  their  lesser  atomic  weights  making 


largely  in  our  favor.  Therefore,  we  will  multiply  their  weight  by 
'2,  and  reckon  them  as  ferrous  oxide.  We  have  then  as  available 
ferrous  oxide: 

•61,067  pounds  iron  in  charge =  78,471    pounds  ferrous  oxide 

18,090  pounds  earthy  bases  X  2 =  36,180 

Total  available  ferrous  oxide. 114,651         " 

This,  with  the  (Jl,410  pounds  silica  in  the  ore,  will  give  a  slag 
containing  34.9  per  cent,  silica.  This  is  somewhat  more  siliceous 
than  we  desire,  nor  have  we  left  any  leeway,  which  is  desir- 
able where  siliceous  ores  abound,  as  it  is  always  extremely  easy  to 
make  the  slag  more  siliceous  if  required,  it  only  being  necessary  to 
select  the  ore  a  little  more  thoroughly,  and  throw  a  trifle  less  work 
•on  the  concentrator. 

We  will  here  leave  the  ore  cupolas  temporarily,  and  take  up  the 
matte  concentration  and  converter  processes,  which  will  show  us 
how  much,  and  what  quality  of  slag  we  shall  have  to  return  to  the 
ore  cupola. 

Matte  Concpntration. — In  this  operation  we  shall  use  a  hot  blast, 
and  shall  depend  for  our  fuel  mainly  upon  the  combustion  of  the 
sulphur  and  iron  in  the  matte  and  added  ore.  Whatever  success 
may  have  attended  pyritic  smelting  in  general,  no  one  who  has 
had  any  experience  in  the  matter  denies  the  ease  with  which  matte 
may  be  thus  concentrated,  providing  that  the  blast  is  heated  to 
800  degrees  to  1,000  degrees  Fahr.  (5148  degrees  C),  that  the  fur- 
nace is  of  the  proper  size  and  shape  to  ensure  sufficient  oxidation, 
that  the  blast  is  low  in  pressure  and  large  in  volume,  and  the 
smelting  is  intelligently  managed.  The  very  small  percentage  of 
coke  that  may  be  used  to  keep  things  in  a  comfortable  condition 
(1  per  cent,  to  3  per  cent.)  will  not  make  ash  enough  to  demand 

The  composition  of  the  charge  will  be  as  follows: 







Matte  from  ore  cupola 

113,000  with 
40.000  with 





3,  (=40 



153,000  with 





37,900  pounds  copper  will  make   68,910   pounds  of  55   per  cent, 
matte,  consisting  of 


Subsulpbide  of  copper 47,375  pounds. 

Sulphide  of  iron 21,535        " 

Total 68, 905 

The  iron  thus  taken  up  in  the  matte  as  ferrous  sulphide 
amounts  to  13,696  pounds,  which,  when  deducted  from  the  total 
amount  of  iron  that  was  in  the  charge,  {52,797  pounds),  leaves 
available  for  slag  formatiou 

Iron 39,101  =  50,245  pounds. 

Earthy  bases,  reduced  to  value  of  ferrous  oxide.  .  3,840  X  2  =    7,680        " 

Total  ferrous  oxide 57,925 

As  there  are  only  15,360  pounds  of  silica  in  this  charge,  the 
above  amount  of  ferrous  oxide  would  make  a  slag  containing  only 
about  21  per  cent,  silica,  which  is  much  too  low  for  good  work. 
As  we  desire  in  this  furnace  to  form  a  slag  containing  about  28 
per  cent,  silica,  we  require  to  add  to  the  charge  about  7,166  pounds 
silica.  It  is  well  to  thus  leave  a  point  where  we  are  sure  of  a  little 
basic  excess,  for  where  siliceous  ores  abound,  the  ore-cupola  slag 
always  seems  to  run  a  little  more  acid  than  we  anticipate,  and  it 
is  easy  enough  to  cancel  this  margin  whenever  desired,  either  by 
using  a  few  more  tons  of  unconceutrated  ore  as  already  suggested, 
or  bv  adding  a  daily  proportion  of  the  rich,  siliceous  slimes  from 
the  concentrator,  which  are  invariably  only  too  plentiful  under 
the  assumed  conditions. 

In  order  to  avoid  too  great  length  of  calculations,  I  will  assume 
that  the  required  silica  is  added  as  pure,  non-cupriferous  quartz, 
although  of  course  tins  would  not  be  done  in  actual  work. 

Then  the  total  weight  of  slag  to  return  from  this  process  to  the 
ore  cupolas  will  be  as  follows: 

Ferrous  oxide 57,925  pounds. 

Silica  in  charge. 15,360         " 

Added  silica 7,166       " 

Total 80,451*     " 

T7ie  converter  plant  will  be  required  to  take  care  of  the  68,910 

♦The  weight  of  the  matte-concentration -cupola  slag  may  be  taken  at  the 
above  figure  in  estimating  its  chemical  effect  in  the  ore  cupolas,  but  its  actual 
weight  is  a  trifle  less,  owing  to  the  doubling  of  the  weight  of  the  earthy  baseSy 
to  make  them  equal  in  effect  to  ferrous  oxide.     This  will  be  corrected  later. 


pounds  of  55  per  ceut.  matte  from  the  matte-concentration  cupola. 
This  matte  has  the  following  composition: 

Copper  '. .  37,900  pounds. 

Iron 18,696 

Sulphur 17,314 

Total 68,910 

The  iron  in  the  above  matte,  13,696  pounds =■  17,600  pounds  FeO. 

To  make  a  slag  of  30  per  cent,  silica  it  will  require 7,543         "       SiOa. 

Total 25,143 

(The  small  amount  of  alumina  taken  up  in  this  slag  from  the 
clay  of  the  converter  linings  may  be  disregarded,  both  because  its 
proportion  to  the  total  amount  of  material  smelted  in  the  ore 
furnaces  is  almost  iniinitesimal,  and  also,  because,  in  such  a  ferrugi- 
nous slag,  its  presence  is  rather  welcome  than  otherwise,  as  tend- 
ing, in  some  slight  measure,  to  decrease  its  specific  gravity.) 

Complete  Ore  Cupola  Charge. — 'We  now  have  the  data  from  which 
to  calculate  the  total  quantity  of  material  that  will  come  to  the 
ore  cupolas,  and  can  thus  estimate  the  quantity  of  coke  required, 
and  allow  for  the  ash  in  the  same.  The  coke  consumption  will  be 
very  low.  Much  of  the  charge  consists  of  sulphides,  and  consid- 
erable heat  will  be  generated  by  the  combustion  of  their  sulphur 
and  iron.  At  the  Butte  &  Boston  smelter  at  Butte,  Montana,  Mr. 
Allen  is  smelting  raw  sulphides  in  a  cupola  with  10  per  cent,  of 
coke,  and  this  without  any  especial  attempt  to  profit  by  the  heat 
resulting  from  their  oxidation.  In  the  case  under  consideration, 
where  the  furnace  will  have  great  area,  perpendicular  walls,  and 
low  pressure  with  large  volume  of  blast,  the  pyritic  effect  will  be 
considerable,  and  both  the  ratio  of  concentration  and  the  con- 
sumption of  carbonaceous  fuel  will  be  benefited  thereby.  But,  for 
conservative  reasons,  I  will  estimate  the  consumption  of  coke  at 
10  per  cent,  on  the  entire  charge,  or  about  13  per  cent,  on  the 
weight  of  the  ore  smelted. 

About  44,0(10  pounds  of  coke  will  be  used,  containing  some  10 
per  cent,  ash,  having  the  following  composition: 

Silica 56  per  cent. 

Ferrous  oxide 21         " 

Earthy  bases 23         " 

Total 100 










219,000  with 

120.000  with 

,HX4ol  with 

25.143  with 

4.195  with 






46  060 

Slag  from  matte  conceutratiou . 
Sla^  from  converters 




444,5W  with 





The  above  fiible  gives  s-iinply  the  siHir-foiniing constituent.s of  the 
ore-cupola  charge  and  the  copper.omittiiigi)art  of  the  oxygen  brought 
into  the  charge  by  the  slags  from  matte-cupola  and  converters, 
and  also  omitting  the  sulphur,  which  possesses  no  interest  for  ns 
in  the  calculation,  and  has  already  been  given  in  the  preliminary 
tignres.  We  can  be  very  certain  that  there  will  be  no  dirticnlty 
in  inducing  the  copper  and  iron  to  take  up  enough  sulphur  to 
form  a  55  per  cent,  matte,  and  all  sulphur  beyond  this  must  be 
burned  to  sulphurous  acid  gas.  This  can  be  done  with  ease  and 
certainty,  and  herein  lies  the  main  chemical  difference  between  the 
old  and  the  most  modern  practice. 

After  deducting  the  4-4,917  pounds  iron,  which,  as  we  found 
before,  is  temporarily  lost  in  forming  the  30  per  cent,  matte,  we 
have  the  following  slag-forming  materials  remaining  in  the  mix- 

Silica 93,943  pounds. 

Eartbs,  reckoned  as  ferrous  oxide 45,884 

Iron,  as  ferrou.s  oxide   147,240       " 

Total 287.067       " 

This  orives  us  a  slair  consisting  of  silica 32.72  per  cent. 

Ferrous  oxide  (or  its  equivalent) 07.28         " 


This  gives  us  our  desired  slag  of  about  33  per  cent,  silica,  with 
such  proportion  of  earthy  bases  as  the  ore  and  coke-ash  will  afford. 
Run  slowly,  and  with  "reverberatory-settler"  (to  be  described 
later),  this  slag  need  not  contain  over  0.4  per  cent,  copper. 

The  actual  weight  of  the  shiw  may  be  determined  by  deducting 
the  amount  of  the  earthy  bases  from  the  total  weight  of  the  ore- 
cupola  slag  already  given  (28?, 067  pounds).  This  is  necessary 
because  the  weight  of  the  earthy  bases  was  doubled,  in  order  to 
make  them    equal   in   chemical   effect   to  ferrous  oxide;  287,007 


.minus  22,942  equals  264,125  pounds,  which  will  be  the  total 
Aveight  of  the  slag  produced  in  the  ore  cupolas  every  24  hours. 

This  is  as  far  as  it  is  suitable  to  carry  the  illustration,  as  it  is 
not  intended,  at  this  point,  to  describe  the  planning  of  the  works 
to  treat  the  above  ore.  I  will  only  add  that  to  smelt  the  222  tons 
of  ore  and  slag  that  are  to  come  daily  to  the  ore  cupolas,  1  should 
use  three  large  blast-furnaces,  tiius  giving  each  a  duty  of  only  74 
tons  of  charge,  or  57  tons  of  ore.  This  low  duty  will  permit  of 
the  slow  running  essential  when  the  blast-furnace  is  to  be  used  as 
an  oxidizer  and  partial  generator  of  its  own  heat,  and  will  also 
permit  ample  stoppages  for  repairs  without  any  diminution  of  the 
ontput.  It  will  scarcely  add  anything  to  the  cost  of  smelting  per 
ton,  as  the  charging  will  be  done  with  mechanical  aid,  and  there 
will  be  one  weighraan  whether  there  are  two  or  three  cupolas. 
As  the  slag  is  to  be  granulated  and  removed  by  water,  the  item  of 
pot-haulers  does  not  enter  into  consideration,  and  the  moderate 
4jnd  comfortable  running  done  by  three  furnaces,  as  compared 
with  trying  to  smelt  the  same  amount  of  ore  in  two  cupolas,  will 
save  enough  work  to  supply  the  labor  needed  below  at  the  third 

The  matte-concentration  will  require  only  one  cupola,  with  hot 
blast.  There  are  only  about  80  tons  of  material  to  treat  daily, 
iind  it  will  be  more  difficult  to  "hold  back"  the  matte,  than  it  will 
to  put  it  through.  It  may  be  necessary  to  run  a  more  siliceous 
slag  at  this  cupola  to  prevent  too  rapid  smelting,  matte  being  so 
heavy  and  so  fusible  that  it  is  difficult  to  restrain  it  long  enough 
to  gain  the  oxidation  necessary  for  its  proper  concentration.  In 
the  present  case  there  will  be  no  trouble,  however,  as  the  concen- 
tration aimed  at  is  less  than  2  into  1.  Any  extra  desired  acidity 
of  the  slag  will  be  a  welcome  circumstance  to  the  concentrator 
foreman,  as  relieving  the  pressure  in  the  slime  department. 

So  little  flue-dust  will  be  made  with  the  light  blast  and  slow 
running,  and  the  capacity  of  the  furnaces  is  so  ample,  that  it  need 
not  be  here  considered. 



The  priucipal  developments  iu  the  American  system  of  blast- 
furnace practice  had  already  long  taken  place  at  the  time  of  the 
publication  of  the  first  edition  of  this  work.  The  improvements 
since  that  time  have  been  characterized  by  perfecting  of  details,  a 
simplification  and  economy  in  the  method  of  manipulating  the 
furnace  and  its  accessory  apparatus,  and  a  decided  saving  in  the 
handling  of  charge  and  product,  rather  than  by  any  radical  change 
of  principles. 

1  do  not  hesitate  to  call  it  the  American  system  of  blast-furnace 
practice;  for  its  advance  on  the  German  process  whence  it  sprang 
is  so  marked,  and  its  whole  style  of  working  so  radically  different, 
as  to  constitute  a  new  departure. 

Twenty-five  years  ago  the  copper  blast-furnace  was  regarded  as 
an  intricate,  eccentric,  and  highly  uncertain  machine,  erected  on 
deep  and  massive  foundations,  enclosed  in  spacious  and  expensive 
buildings,  and  provided  with  one  to  five  tuyeres  of  limited  area, 
through  which  a  gentle  stream  of  air  trickled  into  the  interior, 
Avithout  disturbing  their  most  important  feature,  the  "nose." 

The  tamping-in  the  bottom  of  this  furnace  and  its  long,  brasque 
foreiiearth,  and  its  subsequent  careful  drying,  was  a  ceremony  that 
lasted  days,  and  led  up  to  that  culmination  of  the  metallurgist's 
skill  and  responsibilitv,  the  "  blowing-in."  E\ery  charge  was  then 
watched  as  it  descended,  and  the  subtraction  of  half  a  scoop  of 
coke,  or  the  substitution  of  a  shovel  of  ore  for  a  similar  amount 
of  slag,  were  matters  for  grave  consideration  and  argument.  Even 
after  a  day  or  two  when  the  furnace  was  iu  full  blast, it  was  gener- 
ally thought  necessary  to  use  from  :20  per  cent,  to  50  per  cent,  of 
slag  in  the  cliarge;  and,  indeed,  owing  to  the  imperfect  settling  of 
the  matte,  there  was  seldom  any  lack  of  foul  slag  for  the  purpose. 

Thu  charging  was  done  with  the  utmost  care  and  on  the  most 
minute  scale,  a  charge  often  consisting  of  but  200  pounds,  which 
was  painfully  distributed   around  the  walls  of  the  shaft;  and  the 


slow  smelting,  together  with  the  iufluenoe  of  the  finely-broken 
coke  and  thin  layers  of  ore,  caused  such  a  powerful  reducing  action, 
that  iron  "sows"  were  a  constant  menace  and  frequent  reality. 
Indeed,  certain  smelting  works  were  provided  with«pecial  furnaces,, 
where  these  metallic  masses  were  subjected  to  a  "Verblasen,"  or 
scorification,  to  recover  what  value  they  might  contain. 

The  campaigns  were  short,  and,  like  nations,  were  characterized 
by  a  long  period  of  very  gradual  rise,  a  short  interval  of  maximum 
prosperity,  and  a  protracted  and  most  painful  term  of  decadence 
and  waning  productiveness. 

Fifteen  to  30  tons  of  ore  per  24  hours  was  considered  a  fair  duty 
for  a  copper  furnace,  and  the  campaigns  seldom  lasted  for  more 
than  a  month.* 

The  present  American  copper  cupola  of  the  most  advanced  type 
consists  of  a  circular,  or  oval,  water-jacketed  shell — the  inner  skin 
sometimes  being  of  thick  sheet-copper,  to  withstand  the  corrosive 
action  of  damp  ores  that  contain  sulphates  (quenched  calcines) — 
or  of  four  or  more  straight  wrought  jackets,  that  are  clamped 
together  to  form  the  sides  and  ends  of  a  rectangle,  perhaps  40  by 
160  inches.  The  tuyeres  are  ten  to  twenty  in  number,  contrived 
so  that  their  diameter  may  be  varied,  and  arranged  so  that  the 
blast  in  each  one  may  be  independently  controlled.  The  blast  is 
derived  from  a  positive,  or  semi-positive  blower,  and  furnishes  at 
least  7,000  cubic  feet  of  air  per  minute,  at  a  pressure  of  two  inci}es 
mercury  (090  mm.  water).  The  blast  never  ceases,  except  in  case 
of  accident  or  repairs. 

The  molten  products  escape  at  once  from  the  brick  bottom  of 
the  furnace  into  a  brick-lined  movable  forehearth,  of  large  dimen- 
sions. From  this,  the  thoroughly  settled  slag  flows  in  a  constant 
stream  into  large  pots  drawn  by  mules,  or  into  a  stream  of  water, 
which  granulates  and  removes  it.  In  some  large  works,  the  matte  is 
tapped  in  charges  of  five  tons  into  clay-lined  ladles  moved  by  an 
electric  crane,  which  pours  it  immediately  into  a  Bessemer  con- 
verter, where  it  is  blown  up  to  99  per  cent,  copper  in  a  single 
operation,  and  cast  direct  into  anode-plates  for  electrolytic  treat- 
ment, if  it  contains  the  precious  metals;  otherwise,  into  pigs  for 
the  refinery. 

The  amount  of  foul  slag  to  be  resmelted  need  seldom  reach  one 

*  The  Mansfeld  practice  has  always  been  exceptional,  owing  to  its  unique 


per  cent.,  and  the  substitution  of  a  new  forehearth  every  few  weeks 
is  tlie  only  ordinary  delay;  and  this,  a  very  brief  one.  The  opera- 
tions of  blowing-in  and  blowing-out  are  regarded  about  as  seriously 
as  they  would  be  at  a  foundry -cupola.  In  blowing-in,  the  foreman 
usually  begins  with  a  few  slag-charges,  and  after  a  few  light 
charges  of  ore  the  furnace  is  in  its  normal  working  condition. 
The  charging  is  done  directly  from  cars  or  large  barrows,  and  the 
ore  charge  for  a  furnace  of  this  size  would  be  about  two  tons.  The 
length  of  the  campaign  depends  upon  the  durability  of  the  water- 
jackets  and  machinery,  and  the  prevalence  of  strikes. 

In  a  word,  tbe  American  copper  metallurgist  regards  a  blast- 
furnace as  a  simple  cavity,  surrounded  by  a  fireproof  wall,  in 
which  his  mission  is,  to  burn  coke  witii  the  greatest  attainable 
rapidity,  taking  care  always  to  supply  the  utmost  quantity  of  care- 
fully fluxed  ore  that  the  coke  can  melt,  and  forcing  his  charge 
through  the  furnace  so  quickly  that  there  is  no  opportunity  for  the 
reduction  of  iron  to  a  metal;  whil'^  the  instantaneous  removal  of 
all  molten  material  still  further  prevents  the  formation  of  metallic; 
iron,  enables  tbe  products  of  fusion  to  settle  quietly  and  thoroughly 
according  to  their  weight,  and  removes  the  great  source  of  troubles, 
delays,  and  reiDaiis  from  the  inside  to  the  outside  of  the  furnace. 
A  daily  duty  of  100  to  160  tons  of  ore  is  attained,  and  fron:  late 
experiments  with  ample  blast  and  not  too  fine  ores,  I  have  little 
doubt  that  we  shall  find  it  economical  to  use  furnaces  with  a  daily 
capacity  of  some  300  tons  of  ore. 

The  granulation  of  the  slag  by  water,  and  the  use  of  furnaces 
with  the  gases  drawn  off  below  the  charging-floor,  so  that  the 
tunnel-head  remains  open  and  unobstructed,  is  in  common  use, 
and  permits  the  use  of  an  automatic  car,  the  entire  length  of  the 
furnace,  which  will  drop  its  charge  instantaneously.  (Pueblo 
Silver-Lead  Smelter.)  This  will  remove  any  difficulty  that  might 
be  encountered  in  attempting  to  handle  so  great  a  quantity  of 
material  at  a  single  furnace  in  works  not  suitably  constructed 

The  practice  of  blast-furnace  smelting  in  the  United  States 
almost  invariably  implies  the  employment  of  a  water-jacketed,  or 
water-cooled,  furnace.  Even  the  large  brick  Raschette  furnaces, 
so  skillfully  managed  and  so  firmly  adhered  to  by  the  Orford  Cop- 
per Company,  have  been  cooled  for  many  years  past  by  pipes 
buried  in  the  brick-work  through  which  water  circulates. 

With  so  many  skilled  and   thoughtful  engineers  and  foremen  in 


charge  of  our  copper  plants,  and  in  the  face  of  fhe  grinding  econo- 
mies that  have  nacessarily  accompanied  the  marked  decrease  in  the 
price  of  copper  and  silver,  it  is  not  probable  that  tlie  water-jacket 
would  be  so  universally  employed,  did  it  not  possess  decided  econ- 
omies and  advantages  as  compared  with  its  unprotected  prototype. 
Any  reasonable  suggestion  or  innovation  obtains  patient  hearing 
and  prompt  trial  in  this  country,  and  no  pattern  of  brick  furnace 
that  offers  any  encouragement  for  cheaper  work  would  have,  or 
has  had,  long  to  wait  before  being  somewhere  given  an  opportunity 
to  prove  its  claims. 

During  eiglit  years  of  metallurgical  work  I  used  nothing  but 
uncooled  brick  furnaces,  with,  or  without,  water  tuyeres,  and  1 
think  that  a  brief  comparison  of  general  results  with  subsequent 
water-jacket  practice  may  be  of  interest.  I  feel  the  more  satisfied 
of  the  correctness  of  these  views  from  finding  that  I  hold  them  in 
common  with  all  American  metallurgists  with  whom  I  have  con- 
versed on  the  subject,  ami  u'liose  experience  comprises  both  classes 
of  furnace. 

Where  passable  water  is  obtainable  at  any  reasonable  expense, 
the  first  cost  of  the  two  types  of  furnace  is  pretty  nearly  the  same. 
A  large,  sectional,  copper-lined  water-jacket  of  the  most  modern 
type,  with  deflecting  tuyeres  and  independent  tuyere-valves  will 
cost  considerably  more  than  a  simple,  lightly  built  brick  furnace. 
On  the  other  hand,  a  massive,  thick-walled  brick  furnace  with 
appropriate  foundation  will  cost  more  than  a  plain,  but  perfectly 
good  and  durable  jacket  of  the  Bartlett  type.  And  in  any  case, 
the  difference  in  first  cost  is  but  a  trifling  matter  compared  with 
even  the  slightest  degree  of  efficiency  or  economy  of  one  furnace 
over  the  other.  We  may  assume,  therefore,  the  cost  of  the  two 
furnaces  to  be  equal.  The  main  advantages  claimed  for  the  water- 
jacket  type  are: 

1.  The  -Ease  with  which  it  is  Planned,  Constructed,  and  Erected. 
— It  can  be  planned  at  leisure,  and  the  working  drawings  sent  to. 
the  place  where  it  is  to  be  made.  Then,  after  digging  a  hole  and 
preparing  a  block  of  concrete,  masonry,  or  slag  to  set  it  on,  the 
subject  can  be  dismissed  from  one's  mind  until  the  furnace  arrives 
complete  and  ready  to  set  up. 

It  can  be  erected  by  the  most  ordinary  mechanics  and  in  a  very 
few  days.  This  is  a  great  relief  to  the  metallurgist  accustomed  to 
constructing  brick  furnaces,  with  tlieir  various  items  of  fire-bricV', 
red  brick,  mortar,  clay,  buckstaves,  tie-rods,  arch-patterns,  etc.     I 


have  had  a  water-jacket  furuaoe  rauning  steadily  on  the  third  daj 
after  the  wagons  containing  it  had  arrived. 

2.  The  Simplicity  of  the  BJowing-in  Process. — We  have  learned 
to  make  less  and  less  of  this  once  awe-inspiriug  operation,  yet  even 
now  the  blowing-in  of  a  large  brick  furnace  is  a  slightly  precarious 
task.  The  least  excess  of  fuel  or  pressure  of  blast  is  likely  to 
cause  very  serious  damage  to  the  new  brick-work,  while  an  atom 
too  low  a  temperature  is  certain  to  start  accretions  in  the  hearth 
aud  about  the  tuyeres,  and  too  light  a  blast  may  leave  a  raw  core 
of  ore  in  the  middle  of  the  shaft  that  is  sure  to  cause  much  trouble 
aud  delay. 

This  is  especially  the  case  when  a  new  hearth  or  bottom  has 
been  constructed,  the  proper  drying  aud  warming  of  the  same 
demanding  some  24  hours.  The  heating  up  of  the  great  mass  of 
brick-work  forming  the  shaft  is  also  a  siow  operation,  and  absorhs 
a  vast  amount  of  heat  for  the  first  twelve  hours  or  more. 

But  all  this  is  but  a  small  matter  compared  with  the  burning 
out  of  the  hearth  and  walls.  With  the  fast  driving,  abundant 
coke,  and  basic  charge  usually  employed  in  starting  a  new  furnace 
in  this  country,  it  seems  at  times  impossible  to  maintain  a  perfectly 
uniform  condition  in  the  furnace  shaft,  and  some  one  cornei-  or 
other  is  extremely  likely  to  begin  burning  out,  and  to  defy  every 
etfort  to  stop  it,  until  the  brick-work  is  thinned  enough  to  feel  the 
cooling  influence  of  the  external  air. 

3.  Ease  and  Cheapness  of  Repairs. — This  was  once  a  disputed 
point  between  the  adherents  of  the  two  types  of  furnace.  At  pres- 
ent, the  water-jacket  men  can  find  scarcely  any  one  to  dispute 
with.  The  few  brick  furnaces  that  are  run  in  this  country  are 
managed  with  the  greatest  care  and  skill,  and  every  precaution 
and  manoeuvre  that  years  of  experience  can  suggest  is  brought  to 
bear  npon  them.  With  water-jackets  the  case  is  of  ten  the  reverse. 
These  furnaces  are  constantly  started  at  new  mines  with  inexperi- 
enced men,  and  with  mismanagement  and  abuses  that  are  scarcely 
credible.  The  common  impression  seems  to  be  that  nothing  can 
damage  a  water-cooled  furnace,  and  that  so  long  as  it  does  not 
show  symptoms  of  chilling,  and  the  slag  does  not  carry  too  much 
metal,  everything  is  right.  A  very  small  fraction  of  the  careless- 
ness that  is  so  frequently  displayed  in  running  a  water-jacket 
would  ruin  a  brick  furnace  within  twelve  hours.  It  is  sometin.cs 
a  question  whether  this  extraordinary  capacity  to  withstand  too 
much  fuel,  or  an  improper  slag  or  matte,  is  not  a  positive  disad- 


•vantage,  as  encouraging  waste  and  carelessness.  Consequently,  to 
arrive  at  anything  like  a  fair  comparison  of  the  cost  of  repairs,  we 
must  consider  the  two  furnaces  under  the  same  conditions,  so  far 
as  is  possible.  And,  as  a  brick  furnace  cannot  be  jjrofitably  run 
(in  the  rapid  manner  common  to  this  country)  at  all,  without 
skilled  supervision  and  thoroughly  experienced  foremen  and  work- 
men, we  can  only  compare  it  with  a  water-jacket  run  under  equally 
good  management.  This  greatly  lessens  the  apparent  advantage 
of  the  water-jacket,  as  it  excludes  all  cases  of  careless  management, 
under  which  it  is  probable  that  the  discrepancy  in  the  cost  of 
repairs  and  renewals  would  be  multiplied  many  times  over. 

Under  the  favorable  conditions  referred  to  for  both  furnaces,  1 
estimate  that  the  cost  of  repairs  on  the  brick  furnace,  and  the 
proportion  of  sinking  fund  to  renew  it  when  worn  out,  amount  to 
something  over  double  as  much  as  with  the  water-jacket. 

The  comparative  loss  of  time  from  delays  shows  even  more  un- 
favorably for  the  brick  furnace.  These  points  will  be  considered 
in  detail  when  we  come  to  treat  of  the  expense  of  running. 

4.  Conveuience  of  Maniimlation, — The  advantages  here  are  all 
in  favor  of  the  protected  furnace.  With  the  volume  and  pressure 
of  blast  necessary  to  put  100  tons  or  more  of  ore  through  a  furnace 
every  24  hours,  and  with  the  ordinary  necessity,  or  desirability,  of 
running  a  continuous  stream  of  slag,  there  is  a  strong  tendency 
for  a  portion  of  the  blast  to  escape  through  the  slag-hole,  carrying 
with  it  a  stream  of  glowing  cinders  and  chilled  slag  and  matte- 
globules,  and  often  causing  a  heavy  loss  in  values  on  argentiferous 
ores,  especially  if  a  little  lead,  zinc,  or  antimony  be  present.  The 
loss  in  heat  and  pressure  are  also  considerable,  and  the  workmen 
are  annoyed  by  the  heat  and  noise  of  the  escaping  flame.  In  water- 
jackets,  even  where  there  are  none  of  the  ordinary  water-cooled 
tymps,  or  trapping  devices,  the  escaping  blast  is  suppressed  with 
comparative  ease,  so  long  as  there  exists  a  cold  and  unattackable 
border  to  the  slag-hole,  against  which  brick  and  clay  can  be  solidly 
built.  The  weak  point  of  this  outside  dam  is  its  junction  with 
the  front  wall  of  the  furnace.  In  water-jacketed  furnaces,  as  the 
heated  clay  shrinks  away,  the  resulting  crevice  is  quickly  sealed  by 
the  slag  that  bubbles  out  with  the  blast,  and  a  skilled  furnace-man 
will  make  even  such  a  rude  defense  as  this  last  for  several  hours. 
With  brick  furnaces,  however,  the  crack  widens  rapidly  as  the 
sharp  edges  of  the  brick  work  are  melted  away  by  the  blowpipe 
action  of  the  flame.     More  clay  is  piled  on  in  great  balls  until  the 


front  of  the  furnace  is  provided  with  an  excrescence  resembling  a 
small  haycock,  and  the  buried  brick  front  losing  its  only  chance 
of  being  cooled  (external  radiation),  softens  and  melts  away  more 
or  less  completely,  requiring  the  entire  removal  of  all  the  debris, 
and  the  rebuilding  of  tbe  front.  I  need  hardly  say,  that  in  all 
well-arranged  water-jackets,  the  blast  is  so  trapped  as  to  avoid 
even  the  mild  form  of  loss  and  annoyance  arising  from  the  cause 
just  mentioned. 

Another  highly  important  advantage  possessed  by  water-cooled 
furnaces  is  the  ease  with  which  accretions  are  removed  from  the 
walls  of  the  shaft.  Nearly  all  ores  of  copper  contain  a  little  zinc 
or  lead,  and  a  thick  coating  of  these  metals  soon  settles  on  the 
walls  of  a  furnace.  This  deposit  becomes  so  extensive  in  the 
upjier  portions  of  the  shaft  that  it  would  greatly  lessen  the  capacity 
of  the  furnace  and  also  alter  its  reducing  (or  oxidizing)  power  to 
an  extent  that  would  seriously  affect  the  composition  of  the  matte 
and  slag,  were  it  not  barred  off  from  the  tunnel  head  at  regular 
intervals  of  a  few  days  or  weeks.  The  ore-charge  having  to  be 
allowed  to  sink  as  far  as  practicable,  the  red-hot  walls  of  a  brick 
furnace  make  this  barring  process  a  most  prolonged  and  painful 
task,  not  only  on  account  of  the  excessive  heat,  but  also  because 
the  volatilized  sulphides  soak  into  the  softened  brick-work  until 
they  have  to  be  actually  chiseled  away  at  every  point.  In  the 
water-jacket,  on  the  contrary,  when  the  ore  has  sunk  below  the 
accretions,  the  furnace  shaft  is  com.paratively  cold,  and,  after  a 
charge  of  cold  coke  and  ore  has  been  thrown  upon  the  glowing 
mass  below,  it  is  by  no  means  an  arduous  task  to  bar  away  the 
crusts  from  the  furnace  walls,  especially  as  their  adherence  to  the 
cooled  iron  is  very  slight,  and  when  a  small  portion  of  the  ring  is 
once  chiseled  away,  the  entire  mass  usually  falls  to  pieces. 

In  the  water-jacket  furnace,  the  ojieration  of  blowing  out  is  also 
bereft  of  most  of  its  heat  and  toil,  and  is  so  slight  an  affair  that, 
after  the  charge  has  sunk  nearly  to  the  tuyeres,  the  furnace  can  be 
tapped  dry,  the  forehearth  removed,  the  loose  coke  and  cinders 
still  remaining  dragged  out  and  quenched,  all  within  an  hour,  and 
a  workman  can  immediately  enter  the  furnace  if  repairs  are  neces- 
sary, a  few  inches  of  ashes  being  thrown  onto  the  hearth  to  protect 
the  board  that  he  stands  on.  In  1|  hours  from  the  time  he  is 
through,  slag  can  be  running  again  at  pretty  nearly  the  normal 

The  above  are  a  few  of  the  more  striking  advantages  offered  by 


the  water-jacket  furnace,  but  there  is  a  very  much  longer  list  of 
lesser  advantages  that  will  be  noticed  in  describing  the  mauage- 
ment  of  blast-furnaces,  and  that  form,  when  assembled,  an  over- 
whelming argument  in  favor  of  the  water-jacket  cooled  apparatus. 
I  know  of  but  three  reasonable  arguments  that  are  commonly 
advanced  against  the  employment  of  the  water-jacket.  These 

1.  The  scarcity  and  impurity  of  water  in  certain  localities.  If 
there  is  absolutely  insufficient  water,  it  is  evident  that  a  water- 
jacket  furnace  cannot  be  used.  But  wherever  water  can  be  obtained 
at  any  reasonable  trouble  or  cost,  it  is  equally  certain  that  it  will 
pay  to  do  it.  The  impurity  of  water  has  been  the  cause  of  consid- 
erable annoyance  at  certain  smelters  in  times  past,  but  much  has 
been  done  in  the  way  of  improved  settling  arrangements  for  both 
mechanical  and  chemical  impurities,  and  water-jackets  are  now 
run  steadily  at  places  where  formerly  there  were  many  delays  and 
much  expense  from  this  cause.  Experience  has  also  taught  us 
how  to  construct  the  jackets  to  suit  them  to  such  conditions,  and 
it  must  now  be  a  very  foul  water  that  is  not  preferable  to  no  water 
at  all. 

2.  The  danger  of  ruining  the  furnace  by  careless  management 
of  the  feed  water.  This  is  a  very  curious  objection,  and  applies 
with  much  greater  force  to  steam  boilers  or  to  water-tuyeres  or 
coils.  For  any  overheating  of  the  jacket-water  is  immediately 
shown  by  the  puffing  and  steaming  of  the  discharge  pipe,  and  it 
is  astonishing  how  difficult  it  is  to  seriously  damage  one  of  these 
furnaces,  even  when  there  has  been  the  most  criminal  carelessness 
and  the  jacket  has  been  allowed  to  boil  away  half  its  water  con- 
tents. The  dangerous  temperature  is  all  in  the  neighborhood  of 
the  tuyeres,  and  long  before  the  water  level  has  sunk  to  that  point 
the  furnace  will  have  proclaimed  its  needs  in  a  manner  so  unequiv- 
ocal as  to  startle  even  a  night  foreman  A  mere  stoppage  of  the 
blast  is  sufficient  to  restore  matters  to  comparative  safety  while 
the  fault  is  being  repaired.  And,  lastly,  if  the  furnace-men  are  so 
abnormally  irresponsible  and  unintelligent  as  to  make  it  possible 
that  such  a  condition  of  affairs  should  occur,  it  is  perfectly  easy  to 
arrange  an  alarm  bell  so  that  it  will  act  like  the  danger  signals  on 
the  railroads,  remaining  quiet  while  everything  is  in  proper  condi- 
tion, and  sounding  a  shrill  and  continuous  alarm  as  soon  as  the 
jacket-water  rises  above  a  certain  maximum  temperature.  This  is 
effected  by  an  electrical  connection  with  a  plug  of  fusible  metal  in 


the  water  space,  which  will  melt  at  say  180  degrees  Fahr.  (82  de- 
grees 0.)' 

3.  That  the  water-jacket  wastes  fuel  seriously  in  heatiog  the 
cooliug-water.  This  is  a  grave  charge,  as  much  of  the  blast-furnace 
smelting  in  America  is  done  with  coke  at  $12  to  $15,  and  even  $40 
per  ton.  To  have  any  clear  opinion  on  this  question,  apart  from 
general  knowledge  derived  from  practice  and  comparison,  it  is 
essential  to  tirst  determine  how  much  fuel  is  actually  required  to 
heat  the  water  used  in  cooling  a  jacket  furnace  run  at  the  rapid 
rate  now  generally  adopted.  Mr.  H.  M.  Howe,  in  Bulletin  No.  26 
of  the  United  States  Geological  Survey,  gives  some  figures  on  this 
subject,  made  by  Mr.  J.  B.  F.  Herreshoff,  of  The  Laurel  Hill 
Chemical  Works,  in  1884.  Mr.  Herreshoff  is  such  a  competent  and 
careful  observer  as  to  make  his  figures  of  particular  value. 

The  furnace  was  a  round,  wrought-iron  water-jacket  with  2-inch 
water  space,  the  jacket  extending  from  bottom  of  hearth  to  charg- 
ing door,  and  thus  exposing  an  unusual  area  to  the  heat.  It  was 
52  inches  in  diameter  at  the  tuyeres  and  10  feet  high,  having  ten 
2-inch  tuyeres.  It  averaged  90  tons  (180,000  pounds)  per  24 
hours  of  roasted  6  per  cent,  pyrites,  with  a  consumption  of  12 
tons  of  gas  coke,  making  a  45  per  cent,  matte,  and  a  slag  with  31 
per  cent,  silica,  52  per  cent,  ferrous  oxide,  and  0.55  per  cent, 


Initial  temperature  of  water 15.5  degrees  C. 

Final  temperature . 77 

Gallons  water  per  hour 2,000 

Pounds  of  coke  required  per  24  hours  to  heat  jacket-water,  as- 
suming a  useful  efiFect  of  25  per  cent,  of  the  calorific  power 

of  coke 1,328 

Pounds  coke  for  jacket -water  per  ton  ore  smelted 14.7 

Value  of  this  coke  per  ton  ore  smelted,  at  $5  per  ton  coke $0,039 

Percentage  of  total  coke  consumption  used  in  heating  jacket- 
water 5.5 

I  have  made  a  number  of  similar  tests  on  different  furnaces  run- 
ning on  various  classes  of  ore,  and  under  widely  diverse  conditions 
— although  always  with  large  capacity.  The  results  vary  very 
considerably  according  to  the  state  that  the  furnace  happens  to  be 
in  on  the  day  of  the  test,  and  especially  according  to  the  physical 
condition  of  the  ore — whether  fine  or  coarse,  porous  or  massive, 
wet  or  dry,  etc.     They  are  also  greatly  influenced  by  the  capacity  of 


the  ore  to  form  a  coating  of  lead  or  zinc  sulphides  on  the  inner 
surface  of  the  Jacket,  which  decidedly  lessens  the  loss  of  heat  to 
the  water.  Such  a  coating  is  highly  advantageous  if  it  does  not 
grow  too  rapidly,  and  it  is  preferable  not  to  have  the  rivet-heads 
too  flat  or  countersunk  on  the  interior  of  the  shell,  as  they  give 
just  the  slight  support  required  to  prevent  this  useful  crust  from 
falling  off  at  intervals  into  the  furnace  and  creating  irregularities, 
as  well  as  increasing  the  consumption  of  fuel.  In  the  various  tests 
referred  to,  I  have  found  that  the  coke  wasted  in  heating  the  jacket- 
water  varied  from  2^  per  cent,  to  10^  per  cent,  of  the  total  amount 
used.  I  am  inclined  to  think  that  about  6  per  cent,  is  the  maxi- 
mum allowable  figure  under  normal  conditions  and  that  if  much 
more  than  this  proportion  is  being  used,  one  of  three  things  is 
happening:  Either 
.    1.  Too  much  coke  is  being  charged,  or 

2.  The  method  of  charging  is  wrong,  and  too  much  coke  is  being 
consumed  in  contact  with  the  furnace  walls,  thus  wasting  a  con- 
siderable proportion  of  its  etfect,  or 

3.  The  circulation  of  the  water  in  the  jacket  is  too  rapid,  and 
the  water  is  escaping  too  cold. 

Mr.  Howe's  table  shows  that  the  Laurel  Hill  furnace  is  expending 
about  4  cents  per  ton  of  ore  smelted,  iu^,  heating  its  jacket-water. 
On  90  tons  burden  per  day,  this  is  at  the  rate  of  15  cents  per 

Now  it  is  practically  impossible  to  determine  the  average  loss  of 
heat  by  radiation  from  the  walls  of  a  brick  furnace,  nor  is  there 
the  slightest  sense  or  object  in  comparing  this  factor  with  the  heat 
used  in  warming  the  jacket-water.  If  we  consider  the  question  of 
radiation  at  all,  we  must, compare  the  loss  by  radiation  from  the 
inner  surface  of  the  brick  furnace  with  the  loss  by  radiation  from 
the  outer  surface  of  the  water-jacket,  which,  as  I  need  scarcely 
point  out,  is  largely  in  favor  of  the  latter.  When  wo  come  to  con- 
sider the  fuel  wasted  in  heating  the  jacket-water,  we  can  only 
compare  it  with  the  damage  done  to  the  brick  furnace-walls,  and 
the  heat  wasted  in  raising  them  beyond  a  proper  temperature. 
We  cannot  separate  the  damage  (and  waste  of  fuel)  occasioned  by 
the  heat,  and  that  done  by  the  fluxing  action  of  the  ores  on  the 
fire-brick.  But  it  is  not  in  the  least  necessary  to  make  th'iA  sepa- 
ration,  as  both  these  sources  of  expense  are  avoided  by  water  cool- 


iiig,  and,  consequently,  both  must  be  counted  against  the  brick 
furnace  in  making  our  comparison.* 

Therefore,  if  any  metallurgist  is  not  content  to  pay  15  cents  per 
hour  (at  New  York  prices)  to  guarantee  the  perfect  integrity  of 
his  furnace  walls  and  breast,  he  is  either  more  skillful,  or  more 
ignorant,  than  most  copper  men  in  this  country. 

Commercial  results  and  the  general  testimony  of  skilled,  practical 
metallurgists  are,  after  all,  more  reliable  than  the  imperfect  tests 
and  comparisons  that  we  can  make  on  this  point.  So  far  as  my 
knowledge  extends,  these  are  practically  unanimous  in  regarding 
the  water-jacketed  furnace  as  the  most  convenient  and  economical 
pattern  of  blast  furnace  for  copper  or  lead  ores. 

This  being  the  type  of  furnace  used  almost  exclusively  in  this 
country,  all  general  remarks  on  blast-furnace  smelting  may  be 
considered  to  apply  especially  to  water-jackets.  A  special  section 
will  be  devoted  to  the  consideration  of  brick  blast-furnaces. 


These  may  be  divided  into  two  classes,  according  to  the  material 
of  which  they  are  constructed  : 

1.  Jackets  made  of  cast  iron. 

2.  Jackets  made  of  wrought  iron,  soft  steel,  or  rolled  copper. 

1.  Cast-iron  jackets  are  necessarily  built  in  sections,  tne  various 
jackets  being  assembled  and  clamped  together  to  form  the  complete 
shaft.  The  lead-silver  smelters  have  been  mainly  instrumental  in 
introducing  this  type  of  furnace  into  the  domain  of  copper  metal- 
lurgy. Having  found  it  to  answer  admirably  for  the  quiet,  moder- 
ate, and  regular  furnace  work  characteristic  of  the  well-conducted 
lead-smelting  process,  they  have  naturally  carried  this  furnace  along" 
with  them,  as  the  diminution  of  rich  lead  ores  and  the  transition 
in  depth  from  oxide  to  sulphide  ores  have  forced  them  into  matte 
smelting.  The  rapid  driving  and,  at  times,  fierce  overheat  of  the 
copper  furnace,  accompanied  by  frequent  irregularities  resulting 
therefrom,  and  from  the  less  careful  fluxing  of  the  ores,  make  the 
cast-iron  jacket  inconvenient  for  the  matte  smelter.  I  am  aware 
that  many  excellent  metallurgists  differ  from  me  in  this  opinion, 
but  I  have  run  both  types  of  furnace  under  many  differing  condi- 

*  It  will  be  understood  that  nearly  all  these  remarks  apply  to  the  conditions 
that  prevail  in  America,  where  furnaces  are  usually  run  at  high  pressure, 
smelting  from  75  to  150  tons  per  day. 


tions,  aud,  with  all  reasonable  care  aud  attention,  I  have  found  the 
delays  arising  from  the  occasional  cracking  and  replacing  of  a 
jacket  to  greatly  exceed  any  possible  increased  first  cost  of  the 
wrought-iron  furnace.  Cast  jackets  are  especially  liable  to  crack 
in  cold  climates  and  during  the  operation  of  blowing-in;  and,  as  it 
is  frequently  necessary  for  the  copper  smelter  to  start  up  a  furnace 
for  a  short  campaign,  it  is  of  prime  importance  for  him  to  have  it 
capable  of  withstanding  all  the  fluctuations  of  temperature  that 
may  occur  under  such  circumstances.  I  know  of  no  possible  con- 
ditions under  which  I  should  not  choose  the  wrought-iron  furnace. 
Cast-iron  water-jackets  are  so  thoroughly  illustrated  and  de- 
scribed in  our  modern  text-books  on  lead-silver  smelting,  thac  it 
would  be  contrary  to  the  design  of  this  work  to  repeat  this  infor- 
mation. I  will  only  mention  a  few  practical  points  that  are  of 
especial  importance  to  the  copper  smelter. 

1.  It  is  important  to  obtain  jacket-castings  from  a  foundry  that 
has  had  considerable  experience  in  making  them,  and  has  made  a 
study  of  the  mixture  of  irons  best  suited  to  them. 

2.  A  plan  of  construction  should  be  adopted  that  will  enable  the 
various  sections  composing  the  furnace  shaft  to  be  keyed  together, 
and  unkeyed,  with  the  greatest  possible  facility,  despatch,  and 
firmness.  The  bustle-pipe,  tuyere  branches,  and  feed-water  con- 
nections should  be  as  few  and  simple  as  possible  in  their  arrange- 
ments, and  so  planned  that  they  can  be  taken  down  or  put  up  in  a 
few  moments.  In  this  way,  the  delay  resulting  from  having  to 
change  a  jacket  during  a  campaign  will  be  reduced  to  a  minimum. 

3.  There  should  be  as  few  different  patterns  of  jackets  as  prac- 
ticable in  the  furnace;  otherwise,  too  many  castings  must  be  kept 
in  stock. 

4.  It  prevents  cracking  and  saves  fuel  to  run  the  jacket-water 
pretty  hot,  say  160  degrees  Fahr.  (71  degrees  C),  or  more.  The 
temperature  should  be  kept  as  uniform  as  possible  aud  in  the  vari- 
ous jackets,  though  when  it  is  decided  to  establish  a  continuous 
circulation  through  all  the  jackets  by  connecting  them  externally 
with  2^-inch  U-shaped  bends,  it  greatly  assists  the  circulation  to 
run  the  feed-water  into  the  two  end-jackets,  and  keep  these  20 
degrees  or  30  degrees  cooler  than  the  others. 

The  main  difference  between  running  a  cast-iron  and  a  wrought- 
iron  jacket  is,  that  with  the  former  we  have  to  be  more  careful  in 
war:ping  and  blowing-in  the  furnace,  and  that  we  occasionally  have 
the  unpleasant  and  unprofitable  task  of  changing  a  jacket  while  the 


furnace  is  in  blast.  It  is  a  hot  and  disagreeable  job  at  the  best,  but 
the  foreman  who  takes  the  most  time  and  pains  to  arrange  all  the 
preliminaries,  and  to  make  it  cool  and  comfortable  for  his  men,  will 
usually  be  found  to  accomplish  it  more  thoroughly  and  more 
quickly  than  the  one  who  attempts  to  rush  things  and  to  work  at 
arm's  length  over  a  mass  of  red-hot  ore  and  glowing  coke. 

A  leak  in  a  water-jacket  is  not  often  so  dangerous  a  catastrophe 
as  it  may  seem.  When  it  is  between  the  edges  of  two  adjoining 
jackets,  the  chilled  slag  on  the  inside  will  soon  force  the  water  to 
seek  an  external  path,  and  when  it  is  on  the  outside,  it  can  easily 
be  conducted  away  from  the  base  or  crucible.  Even  when  the  leak 
is  internal  it  may  not  always  be  serious,  as  furnaces  with  an  iron 
base-plate  instead  of  a  brick  hearth  will  usually  right  themselves, 
the  chilled  slag  and  matte  forming  a  dam  between  which  and  the 
walls  of  the  furnace  the  water  will  usually  be  retained  until  it  can 
seek  an  outlet,  perhaps  between  the  bottcm-plate  and  lower  edge 
of  the  jackets.  But  if  the  leak  continues  to  be  serious,  and  cannot 
be  stopped  by  the  use  of  oatmeal,  horse  dung,  and  other  approved 
sediments  introduced  into  the  affected  jacket,  it  is  better  to  change 
the  section  at  once;  for  not  only  is  there  more  or  less  danger  of 
an  explosion,  but  the  hearth  is  suie  to  be  chilled  and  the  flow  of 
slag  and  metal  obstructed  by  the  steam  generated  inside  the  shaft. 
In  a  furnace  with  a  brick  base,  the  water  may  sometimes  be  drawn 
off  from  the  interior  by  driving  a  heavy  steel  bar  from  the  exterior 
to  the  probable  locality  of  the  self-occluded,  interior  pool,  but  it 
is  a  risky  and  temporary  measure. 

If  the  section  of  jacket  is  to  be  changed,  the  foreman  should 
first  make  sure  that  everything  that  is  needed  in  the  operation  is 
at  hand  and  in  readiness  for  immediate  use.  The  charge  being 
pretty  low  in  the  shaft,  its  glowing  surface  should  be  covered  with 
a  thick  layer  of  dampened  small  coke  and  ashes,  and  two  or  three 
long  bars  should  be  driven  down  from  above  close  against  the  inner 
surface  of  the  leaky  jacket,  and  until  their  points  are  in  contact 
with  the  hearth.  The  air  and  water  connections  are  then  quickly 
broken  and  removed,  and  the  fastenings  of  the  condemned  jacket 
are  unkeyed  or  unbolted,  jackscrews,  if  necessary,  being  placed 
against  the  adjoining  sections  to  keep  them  from  being  forced  out 
of  position. 

The  red-hot  ore  and  coke  that  escape  through  the  gap  are 
promptly  dragged  to  one  side  and  quenched  with  water.  The 
glowing  column  of  charge  that  is  seen  on  removing  the  section  of 


jacket  is  ouly  preveuted  from  escaping  en  masse  by  the  bars  driven 
from  above;  aud  opposite  it,  the  heat  is  too  great  to  permit  of  the 
rapid  replacement  of  the  new  section.  By  means  of  a  number  of 
strips  of  heavy,  refuse  sheet  iron,  about  8  inches  wide  aud  some- 
what longer  than  the  breadth  of  the  open  panel,  the  latter  is  tem- 
porarily closed,  the  sheet-iron  strips  being  handled  with  tongs  aud 
bars,  and  inserted  into  the  opening  so  that  they  span  it  from  side 
to  side,  tlieir  extremities  catching  inside  the  two  adjoining  jackets. 
They  are  strengthened  by  short  iron  rods  that  are  also  so  inserted 
as  to  catch  on  the  inner  surface  of  the  jackets.  This  barrier  keeps 
the  glowing  charge  in  place,  but  scarcely  diminishes  the  powerful 
radiation  from  the  opening,  as  the  sheet  iron  becomes  red-hot  in 
a  moment.  One  hundred  pounds,  or  more,  of  well-puddled,  sticky 
clay  is  in  readiness,  and  being  thrown  in  large  balls  against  the 
iron  casing,  flattens  out  aud  forms  a  thick  coating  impervious  to 
the  heat  for  five  or  ten  minutes.  In  this  time,  the  new  jacket 
should  be  replaced,  filled  with  warm  water  from  a  hose,  and  the 
wind  aud  water  connections  made  at  once.  A  light  blast  can  then 
be  put  on,  and  the  furnace  filled  to  its  normal  height.  The  sheet- 
iron  and  clay  that  are  opposite  the  new  section  will  soon  disappear, 
forming,  while  they  last,  a  good  protection  for  the  new  jacket 
until  it  gets  warmed  up  to  its  work. 

The  supporting  of  the  jackets  may  be  effected,  either  from  an 
iron  frame  resting  on  columns,  or  they  may  be  built  up  directly 
on  the  brick  base  of  the  furnace,  or  even  on  an  immovable  base- 
plate. The  former  method  is  the  more  customary  and  convenient, 
as  it  renders  the  hearth  and  shaft  of  the  furnace  entirely  inde- 
pendeut  of  each  other.  The  principal  manufacturers  of  furnaces 
in  this  country  have  various  excellent  designs  that  are  the  outcome 
of  accumulated  experience.  Want  of  space  forbids  my  going  into 
these  details  that  have  been  thoroughly  worked  out  and  established. 

2.  WrougM-Iron  Jackets. — (Also  made  of  soft  steel,  or  rolled 
copper). — This  is  the  most  common  and  useful  type  of  American 
copper  blast-furnaces.  The  simplest  aud  most  general  is  the  ordi- 
nary, circular  wrought-iron  jacket,  extending  in  a  single  piece 
from  below  the  tuyeres  to  a  point  well  up  the  shaft,  the  total 
length  being  usually  from  6  feet  to  9  feet.  The  diameter  usually 
increases  toward  the  top  at  the  rate  of  one  inch,  or  more,  per  foot 
of  vertical  length.  The  tuyere  openings  consist  of  cast  rings  in- 
serted into  the  water-space,  a  circle  o^  rivets  holding  the  inner  and 
outer  shells  in  close  contact  with  these  castings.     The  water-space 



at  top  and  bottom,  and  the  slag  opening,  may  be  closed  bv  wronght- 
iron  rings  to  which  the  two  shells  are  riveted,  or  the  inner  shell 
may  be  flanged  over  and  riveted  to  the  outer,  forming  a  right  angle 
to  the  vertical  axis  of  the  furnace.  The  bottom  may  consist  of 
drop-doors  like  a  fonndry  cupola,  or  of  a  simple  cast-iron  plate 
(dished,  to  prevent  cracking),  and  bolted  to  the  bottom  ring  of  the 

Fig.  26. 

furnace;  or  it  may  be  built  up  from  the  ground  in  the  shape  of  a 
brick  hearth  of  circular,  oval,  or  rectangular  form,  securely  braced 
by  an  iron  shell.  The  accompanying  illustration  of  Bartlett's 
water-jacket  (Fig.  2G)  shows  this  simple  and  economical  type  of 
furnace.  In  large  smelters,  where  greater  capacity  is  required, 
the  circular  shape  of  furnace  cannot  be  adopted,  as  a  moderate 
blast  will  not  penetrate  a  column  of  average  charge  to  a  greater 
depth   than  20  inches  to  ii  inches.     Consequently,  a  diameter  of 



48  inches  or  50  iuohes  at  the  tuyeres  seems  to  rae  the  extreme 
limit  in  this  direction,  even  with  a  coarse  charge.  By  allowing 
the  tuyeres  to  penetrate  the  furnace  shaft  a  few  inches,  the  diame- 

Figs.  38. 

ter  of  a  circular  shaft  may  be  increased  to  54  inches,  or  even  to  60 
inches,  with  a  nearly  corresponding  increase  of  capacity;  but  the 
complications  resulting  from  the  necessity  of  cooling  these  pro- 
jectiii<i  tuveres,  and  from  otlier  causes,  have  thus  far  outweighed 
the  advantage  gained. 

26d         ■  MODERN    COPPEK    SMELTING. 

Obviously  our  only  recourse  is  to  lengthen  the  furnace  shaft  in 
one  iHrection,  keeping  it  sufficiently  narrow  in  the  other  dimen- 
sion for  the  blast  to  penetrate  to  the  center.  This  brings  us  at 
once  to  the  rectangular  form,  or,  if  we  desire  to  still  make  the 
entire  jacket  in  one  piece,  we  may  construct  it  in  the  shape  of  a 
flattened  oval.  Mr.  J.  B.  F.  Herreshoff  of  the  Laurel  Hill  Chem- 
ical Works,  New  York,  has  done  this  with  much  success,  his 
improved  furnace  being  shown  in  Figs.  27,  28. 

*Figs.  29,  30,  31  and  32,  are  illustrations  of  water-jacket  fur- 
naces largely  employed  throughout  the  West  in  producing  black 
copper  from  ox}dized  ores,  or  matte  of  tolerably  high  grade  from 
sulphide  ores.  It  is  evident  that  the  hearth  would  not  stand  very 
large  amounts  of  low  grade,  fiery  matte  (below  30  per  cent,  cop- 
per); but,  for  the  purposes  intended,  the  cooling  by  radiation  is 
generally  sufficient  to  keep  the  bottom  cool,  though  Walker  has 
lately  found  it  useful  to  use  this  radiated  caloric  in  preheating  the 
blast,  at  the  same  time  keeping  the  hearth  at  a  safer  temperature. 
One  to  two  thousand  gallons  of  water  per  hour  (3,785  to  7,570 
liters)  is  needed  to  cool  the  jacket.  The  water  is  admitted  through 
the  pipes  F,  and  escapes  through  G.  Hand-holes  E  are  very 
essential,  as  the  integrity  and  life  of  the  jacket  depend  largely 
npon  the  care  that  is  given  to  keep  its  interior  free  from  mud  and 
lime-scale.  The  hearth  M  rests  upon  th.e  drop-bottom  P,  and  is 
built  up  of  fire-brick  and  clay.  The  slag-notch  is  at  L,  and  the 
tap-hole  for  the  metal  at  0.  The  entire  furnace  rests  on  the  four 
short  columns  R,  and  is  covered  by  the  hood  H  leading  to  the 
stack  K.  A.  furnace  of  this  description,  42  to  46  inches  in  diame- 
ter at  the  tuyeres  and  6  to  9  feet  high  from  tuyeres  to  charge  door, 
smelts  from  40  to  80  tons  of  ore  per  day.  It  is  usually  driven  by 
a  No.  4^  Baker  blower,  running  100  to  120  revolutions,  and  fur- 
nishing some  2,000  cubic  feet  of  blast  per  minute. 

In  rectangular  wrought-iron  jackets,  the  shaft  may  be  divided 
into  narrow  sections,  as  with  cast-iron  jackets,  or  each  side  and 
end  of  the  shaft  may  be  formed  by  a  single  jacket.  Figs.  33  and 
34  show  this  latter  form  of  construction,  though  in  this  case  there 
are  two  tiers  of  jackets,  one  above  the  other.  The  rectangle  is  32 
inches  by  72  inches  at  the  tuyeres  inside.  The  upper  jackets  B 
are  supported  by  the  columns  L,  while  the  lower  jackets  A  rest  on 

*  Figs.  29  to  34,  with  a  portion  of  the  accompanying  descriptions,  are  taken 
from  the  valuable  paper  of  A.  F.  Wendt.  in  Vol.  XV.  of  the  Transactions 
American  Institute  Mining  Engineers. 

Fig.  32. 



the  bed-plate  carried  by  the  posts  K.  HU  are  the  tapholes  and  M 
the  slag  notches.  The  upper  part  of  the  furnace  is  surrounded  by 
the  shell  0,  and  contains  a  charging-bell  and  hopper  which  is 
worked  by  the  levers  V.     This  has   been  replaced   by  a  simple 

Fig.  3;1 

hopper.  The  capacity  of  such  a  furnace  varies  so  completely  with 
the  ores  and  blast  nsed,  that  it  is  impossible  to  speak  of  it  accu- 
rately except  for  known  conditions.  It  can  smelt  from  60  to  110 
tons  ore  per  24  hours. 



Two  main  objections  are  occasionally  found  to  these  large  rec- 
tangular wrought-iron,  or  soft  steel  water-jacketed  cupolas.  These 
are  the  corrosion  of  the  upper  portions  of  the  jacket  by  material 
containing  sulphate  of  copper,  and  the  buckling  or  distortion  of 

Fig.  34. 

the  inner  shell  of  the  large  jackets  that  form  the  long  sides  of  the 
rectangle.  Herresholf  has  long  substituted  copper  for  iron  for  the 
inner  shell,  to  obviate  the  first  difficulty,  while  the  second  is  over- 
come by  dividing  up  the  long  sides  of  the  furnace  into  several  sec- 


tions.  This  has  nowhere  been  more  perfectly  done  than  in  a  late 
furnace  erected  at  W.  A.  Clarke's  Verde  mine,  near  Prescott, 
Arizona.  The  original  idea  of  the  furnace  was  given  by  Mr.  J. 
L.  Giroux,  the  details  being  worked  out  by  Fraser  &  Chalmers, 
whose  long  experieuca  in  snch  work  has  enabled  them  to  steer 
clear  of  the  difficulties  so  often  encountered  in  new  designs.  Plate 
IX.  shows  this  furnace  in  detail. 

It  consists  of  three  tiers  of  sectional  water-jackets,  extending  from 
the  cast-iron  base-piate  to  the  charging  door,  wliich  is  9  feet  above 
the  tuyeres.  The  middle  tier  of  jackets  has  a  bosh  near  the  bot- 
tom, and  the  upper  tier  is  so  set  that  it  tumbles  in  toward  the 
tunnel  head.  Thus  the  side-walls  are  contracted  at  the  hearth 
and  at  the  tunnel  head,  and  widen  out  at  the  middle  of  the  shaft. 
The  end  walls  are  vertical. 

The  inner  shell  of  all  the  jackets  is  made  of  |-inch  copper,  and 
the  outside  shell,  of  ^j-hvA\  flanged  steel,  the  two  shells  being  stif- 
fened by  stay  bolts  that  pass  across,  tnrough  the  water-space. 
These  stay  bolts  have  caused  no  leakages.  There  is  a  partitioned 
wind-box  containing  10  tuyere  openings,  the  castings  being  of 
phosphor  bronze.  Each  opening  is  provided  with  a  deflecting 
nozzle,  and  a  ball-valve  to  control  the  blast. 

The  inside  dimensions  of  this  furnace  are: 

Length 90  inches. 

Width  at  tuyeres 36        " 

Width  5  feet  above  tuyeres 61        " 

Width  at  tunnel  head 48        " 

Width  of  water  space  in  upper  tier  of  jackets 4        " 

"  "         "      in  two  lower  tiers 5^      " 

Total   weight  of  jackets,  base-plates,  I-beams  for  supporting 

brick-work,  etc 24,000  pounds. 

Total  copper  in  furnace 10,181       " 

Unless  a  metallurgist  has  had  long  experience  with  the  various 
forms  of  water-jacket  and  the  various  details  belonging  to  them, 
and  unless  he  knows  exactly  what  he  is  doing  and  why  he  is  doing 
it,  it  is  far  safer  to  trust  to  the  established  manufacturers  of  this 
apparatus,  than  to  attempt  to  originate  any  improvements  that 
diverge  very  radically  from  the  regular  type.  Inventing  is  one  of 
the  most  expensive  amusements  belonging  to  metallurgy,  and 
should  be  generally  left  to  those  individuals  who  are  led  toward  it 
by  experience  and  talent. 

Having  obtained  and  erected  the  furnaces  and  blowing  plant, 


CMS    rt.fl^A'riCIV. 



there  remains  only  to  make  the  water  and  blast-connections,  put 
in  and  dry  the  furnace  bottom,  and  prepare  and  heat  the  forehearth. 
Care  shonld  be  taken  in,  planning  the  furnace,  that  there  are  no 
narrow  passages  or  pockets  in  the  water-space,  where  sediment  and 
scale  can  collect.  Otherwise,  these  will  quickly  block  up  and  burn 
out.  Such  places  may  be  between  the  tuyere-castings  and  the 
lower  ring  of  the  jacket,  or  between  the  slag  notch  and  lower  ring. 
If  any  such  exist,  it  is  an  excellent  plan  to  drill  a  small  hole  from 
below  into  the  narrow  spot  and  put  in  a  one-half,  or  three-quarter- 
inch  pipe,  through  which  a  constant  current  of  feed  w-ater  forces 
itself  upward  under  pressure,  and  thus  prevents  the  collection  of 
sediment  or  scale. .  This  does  not  supersede  the  frequent  opening 
of  the  hand-holes  and  inspection  of  the  water-space  for  sediment  or 
scale,  and  a  frequent  blowing-oif  of  the  jacket  under  all  the  pres- 
sure possible,  or  assisted  by  the  introduction  of  live  steam. 

Water-jacket  furnaces  have  two  classes  of  bottoms.  In  one 
class,  the  bottom  is  made  as  in  Fig.  30,  being  a  thick  mass  of  brick 
and  clay,  and  dependent  for  its  integrity  on  the  comparatively 
high  grade,  of  the  product  made.  As  every  metallurgist  knows, 
rich  matte,  or  metallic  copper,  tends  to  fill  up  a  hearth  rather 
than  cut  it  out,  and  in  a  furnace  producing  such  material,  there 
is  no  need  of  having  the  water-cooling  extend  down  to  the  bottom 
of  the  hearth,  or  even  to  provide  a  water-cooled  slag-notch,  save 
under  exceptional  circumstances.  Where  the  product  is  a  matte 
of  lower  grade  and  especially  if  made  in  considerable  quantity, 
such  a  bottom  as  the  one  shown  in  Fig.  30  would  soon  be  cut 
through  and  the  hearth  destroyed.  By  extending  the  water-jacket 
down  to  the  bottom  yjlate,  the  sides  of  the  hearth  are  rendered 
safe,  but  the  bottom  is  still  vulnerable,  and,  in  fact,  in  running  at 
high  speed  on  a  matte  of  35  per  cent,  copper  or  less,  the  bottom 
plate  would  be  eaten  through  and  the  contents  of  the  hearth  would 
escape  within  an  hour  or  two.  This  catastrophe  only  takes  place 
from  the  cutting-down  action  of  the  slag  and  matte  at  the  slag- 
notch;  and  if  this  notch  be  jacketed  all  around  so  that  its  level 
cannot  be  changed,  as  in  the  HerreshofE  furnace,  Fig.  27,  the 
hearth  of  the  furnace  up  to  the  lower  edge  of  the  slag  and  matte- 
notch  G,  will  always  contain  a  pool  of  stagnant  metal  that  is 
scarcely  affected  by  the  hot  products  resting  upon,  and  flowing 
over,  its  surface.  The  bottom  of  this  quiet  pool  of  matte  is  in 
contact  with  tlie  thin  layer  of  fire-brick  which  alone  separates  it 
from  the  cast-iron  bottom-])late  E,  and  thus  radiates  heat  so  rapidly 


that  it  soon  becomes  chilled  and  practically  forms  the  bottom 
proper  of  the  furnace.  If  the  furnace  is  run  hotter  or  faster,  or 
with  a  greater  proportion  of  matte,  or  on  a  matte  of  lower  grade, 
an  inch  or  two  of  the  surface  of  this  artificial  bottom  will  be  cut 
away,  and  this  will  continue  until  the  radiation  of  heat  through 
the  thinned  bottom  exactly  balances  the  accession  of  heat  from 
the  smelting,  when  it  will  again  become  stationary.  This  building- 
np  and  cutting-down  of  the  bottom  is  entirely  automatic  and  re- 
quires no  attention  or  assistance  from  the  metallurgist.* 

The  drying  of  the  thick  bottom,  as  in  Fig.  30,  may  require  18 
hours  or  more,  as  it  is  undesirable  to  leave  any  moisture  to  form 
steam.  Where  this  occurs,  a  boiling  action  of  the  molten  products 
is  set  np  that  is  apt  to  result  in  loosening,  or  partially  destroying 
the  bottom. 

The  thin  bottom,  as  shown  in  Fig.  27,  is  usually  dried  for  a  few 
hours,  a  small  wood  or  coke  fire  being  maintained  in  it,  and  the 
ashes  removed  from  time  to  time  that  they  may  not  form  a  non- 
conducting layer  between  the  heat  and  the  bottom.  But  even  this 
slight  drying  is  hardly  essential,  as  the  object  of  the  single  layer 
of  brick  that  forms  the  bottom  is  merely  to  keep  the  hot  metal 
away  from  the  bottom  plate  until  an  artificial  bottom  is  built  up 
by  the  chilled  matte  and  slag. 

In  smelting  oxidized  ores  for  black  copper,  the  bottom  is  made 
deep  enough  to  form  a  small  crucible  for  the  accumulation  of  the 
metallic  product,  which  has  too  high  a  melting  point  to  attempt 
to  collect  in  an  unheated  outside  forehearth.  With  this  exception, 
however,  it  is  the  ordinary  practice  in  the  United  States  to  allow 
no  accumulation  of  molten  material  inside  the  furnace,  but  to 
etfect  the  separation  of  matte  and  slag  in  an  independent  outside 
forehearth,  or  well.  Next  to  the  introduction  of  the  water-jacket 
furnace,  I  regard  this  practice  of  universally  settling  the  matte 
outside  of  the  furnace,  and  thus  removing  all  material  as  soon  as 
possible  from  the  hearth,  as  the  most  important  advance  of  this 
generation  in  the  blast  furnace  treatment  of  copper. 

It  may  be  assumed  that  the  ordinary  diflSculties  experienced  in 
running  a  furnace  with  brick  hearth  built  up  from  the  ground  and 
with  interior  crncible  or  sump,  are  mainly  due  to  its  filling  up 

*  This  principle  f)f  automatic  regulation  by  radiation  has  a  wide  practical 
bearing  in  metallurgrical  operations.  It  is  also  a  good  example  of  the 
advantage  of  accoraplishiner  an  object  by  enlisting  natural  forces  in  our  bebalf 
instead  of  struggling  to  oppose  them. 


with  stick}',  half-fused  products  that  become  more  and  more  diffi- 
cult of  removal,  and  finally  accumulate  until  the  furnace  must  be 
blown  out. 

I  need  hardly  consider  the  opposite  condition  of  affairs  where 
the  hearth  is  cut  awily  and  deepened,  until,  in  some  of  our  large 
brick  furnaces  it  may  contain  25  tons,  or  more,  of  matte.  This 
occurs  only  when  produciug  large  quantities  of  very  low-grade 
matte,  8  per  cent,  to  15  per  cent,  copper,  and  usually  happens 
during  the  reducing  smelting  of  raw  pyrites  fines.  If  the  hearth 
and  fouiidatious  of  the  furnace  are  properly  constructed,  it  is  best 
to  let  matters  take  their  course,  feeling  sure  that  when  the  matte 
has  cut  its  way  down  deep  'enough  to  make  the  radiation  below 
equal  to  the  accessions  of  heat  from  above,  it  will  cease  burrowing 
of  its  own  accord.*  This  leaves  a  permanent  bottom,  containing 
perhaps  40  tons  of  15  per  cent,  matte,  or  12,000  pounds  copper, 
worth  perhaps  7  cents  per  pound  in  this  condition,  or  $840  in  all. 
At  6  per  cent,  per  annum,  this  amounts  to  14  cents  per  day,  or 
about  one-seventh  of  a  cent  per  ton  of  ore  smelted,  which  is  not 
an  extravagant  price  to  pay  for  the  luxury  of  a  bottom  that  re- 
quires neither  renewals  nor  repairs. 

Hence,  we  may  fix  our  attention  on  the  filling-iip  rather  than 
the  cutiing-down  of  the  crucible.  While  the  accretions  that  so 
frequently  form  in  the  hearth  of  a  furnace  with  interior  crucible 
are  often  termed  sows,  salamanders,  or  bears,  it  is  seldom  that 
they  are  entited  to  these  designations,  which  are  more  correctly 

*  The  principle  of  automatic  regulation  by  radiation  is  again  illustrated  in 
tbis  practice.  With  competent  and  experienced  furnace-men  there  is  scarcely 
a  limit  to  the  time  which  such  a  bottom  will  last,  being  constantly  torn  down 
and  built  up  by  its  own  internal  processes.  It  is  the  furnace-men's  duty  to 
assist  these  matters  by  various  well-known  means  at  their  disposal,  among 
which  the  commonest  are: 

Using  an  excess  of  pyrites  and  a  heavy  blast,  so  as  to  make  a  poorer  matte 
and  cut  down  a  bottom  that  has  grown  too  high. 

Using  less  pressure  of  blast,  but  larger  tuyeres,  to  effect  a  more  forcible 
oxidation  of  the  pyrites,  and  thus  make  a  thin,  ferruginous  slag,  and  a  richer 
and  scantier  matte,  which  will  soon  build  up  a  vanished  bottom.  Or,  if  the 
bottom  seems  to  be  cutting  down  beyond  all  bounds,  allowing  the  furnace  to 
stand  without  blast  for  several  hours,  during  which  time  radiation  from  the 
crucible  will  be  going  on  without  any  accession  of  heat  from  above.  This  is  a 
very  certain  means,  and  will  soon  lay  the  foundations  of  a  solid  hearth,  which 
is  built  up  still  more  by  the  richer  and  more  infusible  matte  produced  when 
the  blast  is  again  let  on  to  the  charge  which  has  been  slowly  roasting  during 
the  period  of  repose. 


applied  to  accretious  cousisting  maiulv  of  metallic  iron.*  These 
are  geuerally  of  gradual  growth  and  are  produced  most  freely  in 
furnaces  where  the  smelting  is  slow  in  comparison  with  the  hearth 
area  where  there  is  a  high  ore  column,  a  contracted  hearth 
(boshes),  and  a  scarcitij  of  iron  or  other  bases.  Paradoxical  as  it  may 
seem  at  first  glance,  a  scarcity  of  iron  (or  proper  bases)  in  the  slag 
causes  iron  in  metallic  form  to  be  separated  out  from  this  same 
slag.  Yet  the  reason  is  qaite  obvious.  By  withdrawing  iron  from 
the  slag,  we  decrease  its  fusibility  and  raise  its  smelting  point. 
Now  a  siliceous  slag  with  high  melting  point,  produced  in  a  furnace 
intended  for  a  more  fusible  mixture,  brings  about  slow  smelting, 
a  rising  of  the  heat  toward  the  tunnel  head,  a  powerful  reducing 
action,  and,  in  a  word,  inaugurates  on  a  small  scale  the  condition 
of  affairs  prevailing  in  furnaces  devoted  to  the  production  of  pig 
iron  from  ores.  We  have  not  a  sufficiently  high  temperature  nor 
reducing  power  to  form  the  ferrous  carbide  that  we  know  as  cast 
iron,  but  we  can  produce  an  infusible  wrought  iron  with  the 
greatest  facility. 

Such  conditions  are  rare  in  America,  as  rapid  driving  and  the  pro- 
duction of  fusible  slags  by  the  avoidance,  or  mechanical  concentration 
of  siliceous  or  aluminous  ores  are  opposed  to  the  formation  of  metal- 
lic sows.  Hence,  the  accretions  that  we  find  in  our  furnaces  are 
apt  to  be  mixtures  of  magnetic  oxide  of  iron  with  infusible  slags, 
indefinite  compounds  of  baryta,  zinc  oxide,  etc.,  with  which  is 
usually  interspersed  a  quantity  of  very  basic  matte;  that  is  to  say, 
a  matte  with  too  large  a  proportion  of  iron  and  too  little  snlpht.r. 
These  accretions  are  very  hard  to  break  up,  even  when  outside  the 
furnace,  and  are  so  slippery  and  intangible  when  at  a  high  tem- 
perature, that  it  is  very  rlifficult  to  drag  them  out  of  the  interior. 
We  escape  this  filling  up  of  the  furnace,  and  the  serious  labor  and 
delays  attendant  thereon,  by  transferring  the  settling  process  to  an 
outside  crucible  (forehearth  or  well),  which  is  not  only  accessible, 

*  A  sample  of  borings  from  such  a  chill,  analyzed  for  the  writer  by  Mr.  A.  F. 
Glover,  Ph.D.,  had  the  following  composition: 

Sulphur 4.64 

Copper 9.80 

Iron 82.70 

Carbon 1.13 

Arseuic 0.41 

Slag 0.78 

Nickel  and  cobalt 0.81 



and  thus  easy  to  clean  out  and  repair,  but  which  can  be  removed 
and  replaced  inside  of  an  hour.* 

The  four  main  causes  of  trouble  and  delay  in  the  running  of 
copper  blast-furnaces,  and  the  means  generally  adopted  in  the 
United  States  for  the  avoidance  of  these  inconveniences,  are: 


1.  Destruction  of  lining.  Water-cooled  walls. 

2.  Choking-up  of  shaft  by  accretions.      Metallic    water-cooled    shaft,    which 

permits  their  easy  removal. 

3.  Burning-out   of   crucible   and  bot-      Self-created      bottom,     automatically 

torn.  regulated  by  radiation. 

4.  Filing  up  of  crucible  or  hearth  by      Permitting  all  molten  material  to  run 

sows,  or  other  accretions.  out  of  hearth  as  soon  as  it  can,  and 

thus  transferring  possible  accre- 
tions to  an  external  and  exchange- 
able forehearth. 

The  blowing-in  of  a  modern  water-jacket  copper  furnace,  on 
known  ores,  whether  small  or  large,  would  be  scarcely  worth 
alluding  to,  were  it  not  that  traces  of  the  anxiety  and  importance 
that  once  attached  to  this  operation  still  hang  about  it. 

In  starting  a  brick  furnace  there  is  a  large  mass  of  material  to 
be  warmed  up,  and,  above  all,  most  of  this  material  can  be  de- 
stroyed by  an  excess  of  heat,  or  by  a  very  trifling  want  of  skill  in 
fluxing  or  management.  In  a  water-jacket,  however,  the  only 
extra  caloric  required  for  warming  up  is  tlie  few  heat  units  neces- 
sary to  raise  the  jacket-water  to  its  normal  temperature,  and  to 
prepare  the  cold  bottom  for  the  molten  matte  and  slag  that  are 
soon  to  cover  it.  But  the  feed-water  and  bottom  have  probably  both 
been  heated  up  by  a  preliminary  drying  fire,  and  a  few  inches  of 
bot,  low-grade  matte  will  do  more  to  get  the  bottom  in  proper 
condition  than  hundreds  of  pounds  of  coke.  Hence,  in  blowing-in 
we  have  no  use  for  any  extra  fuel  except  to  heat  the  first  few  thou- 
sand pounds  of  slag  and  matte  sufficiently  beyond  their  proper 
normal  temperature  to  provide  enough  heat  to  warm  the  bottom, 
and  especially  the  external  forehearth,  up  to  its  regular  condition. 
As  the  forehearth  has  already  been  brought  up  to  a  red  heat  by 
means  of  a  bushel  or  two  of  coke  or  charcoal  (aided,  perhaps,  by  a 
light  blast  through  the  tap-hole  at  the  side),  the  amount  of  heat 
to  be  abstracted  from  the  molten  products  to  bridge  the  space  be- 
tween red-heat  and  the  normal  white-hot  condition  of  the  forehearth 

*The  subject  of  forehearths  is  sufficiently  important  to  demand  a  separate 
section  for  its  consideration. 


is  very  small.  Beyond  the  fuel  necessary  to  supply  this  slight 
amount  of  heat,  every  pound  of  extra  coke  is  a  positive  and  serious 
detriment  in  various  directions.  The  two  most  obvious  evils  arn: 
The  waste  of  money  in  consuming  coke  to  uselessly  heat  the 
jacket-water,  and  the  much  more  serious  matter  of  reducing  iron 
out  of  the  slag  by  the  high  temperature  and  powerful  reducing 
action  arising  from  the  excess  of  fuel. 

This  over  use  of  .coke  in  blowing-in  a  water-jacket  (where  one 
loses  the  wholesome  restraint  imposed  by  the  fear  of  damaging  the 
lining),  is  so  common  and  serious  an  error  that  it  seems  worth 
while  to  illustrate  it  by  an  example. 

Some  years  ago  I  was  present  at  the  starting  of  a  large  water- 
jacket  furnace  in  Southern  Arizona.  The  ores  were  pure  carbon- 
ates and  oxides,  and  tl)e  slag  was  to  be  rather  siliceous  and  low  in 
iron;  lime  and  alumina  being  more  accessible  than  were  ferrugi- 
nous ores.  The  slag  had  been  carefully  calculated,  and  appeared 
to  be  a  feasible  one,  though  the  former  owners  of  the  mines  had 
run  a  highly  ferruginous  slag,  exhausting  all  the  cupriferous  hem- 
atite that  could  be  found  in  the  neighborhood.  The  furnace  had 
been  in  hlast  six  hours  when  I  first  saw  it,  and  presented  a  very 
sickly  appearance.  The  slag  was  only  red-hot  and  very  scanty, 
and,  apparently,  extremely  siliceous.  No  copper  could  be  found 
on  trying  the  tap-hole,  and  all  10  tuyeres  had  long  noses  that 
united  at  the  center  of  the  shaft,  and  through  which  not  the 
feeblest  glow  of  heat  could  be  seen.  The  charge  sank  extremely 
slowly  and  irregularly,  the  jacket-water  was  almost  boiling  in  spite 
of  a  full  supply  through  an  ample  feed-pipe,  and  the  heat  was 
mounting  to  the  tunnel  head.  There  were  obviously  strong  pru- 
dential reasons  against  blowing  out  and  starting  afresh. 

As  is  usual  in  such  cases,  the  furnace  had  been  started  with  an 
enormous  excess  of  coke,  and  even  after  six  hours  running,  only  one- 
lialf  the  charge  that  this  coke  was  expected  to  support  had  been 
reached.  Yet  the  furnace-men  were  clamoring  for  a  few  "empty 
charges"  (coke  without  ore)  in  order  to  heat  up  the  slag  and  melt 
out  the  solid  mass  that  pretty  nearly  filled  the  hearth  and  lower 
portion  of  the  shaft.  If  the  furnace  had  been  a  small  one  it  could 
scarcely  have  been  saved,  but  it  requires  a  considerable  amount  of 
time,  as  well  as  metallurgical  skill,  to  completely  freeze  up  one  of 
the  long,  rectangular  shafts  now  in  such  common  use,  and  there 
was  still  hope. 

A   totally  new   departure   was   agreed    upon.     Every  alternate 


•tnyere  was  phiggetl,  the  blast  was  reduced  to  a  quarter  of  an  iucli 
•of  mercury,  the  coke  charge  was  maintained  at  300  pounds,  but 
instead  of  small  charges  of  ore  they  began  on  3,000-pound  charges 
of  the  old  (rich)  ferruginous  slags.  These  were  continued  for  five 
charges,  when  one-half  of  the  slag  was  replaced  by  ore,  and  later, 
ore  was  gradually  substituted  for  the  remaining  1,500  pounds  of 
slag  at  the  rate  of  40  pounds  ore  for  100  pounds  slag,  so  that  the  nor- 
mal charge  became  2,100  pounds  ore  to  300  pounds  coke,  with  the 
addition  of  the  2  or  3  per  cent,  of  foul  slag  made  by  the  furnace. 
For  some  two  hours  after  the  change,  things  looked  very  bad. 
The  slag  stojDped  running  almost  completely,  and  the  wind  blew 
through  the  cold  noses  of  the  tuyeres  as  though  its  only  effect 
■were  to  remove  what  little  warmth  still  remained  in  the  siliceous 
skeleton  inside.  But  finally  the  plugged  tuyeres  began  to  brighten 
one  by  one,  and  then  it  was  evidently  only  a  question  of  time.  It 
is  almost  as  hard  to  damage  an  improving  furnace  as  it  is  to  better 
a  sickly  one.  The  bright  tuyeres  were  put  in  blast  as  they  became 
fit  for  it,  and  their  chilled  neighbors  were  plugged  in  turn,  nntil 
eventually  there  were  no  noses  left,  and  as  the  heavy  charges  of 
basic  slag  came  down  and  swept  the  siliceous  chill  before  it,  the 
furnace  went  to  the  other  extreme,  and  it  was  impossible  to  handle 
the  slag  with  the  12  pots  provided.  Some  holes  were  dug  to  one 
side  in  the  floor,  and  many  tons  of  slag  run  into  them,  to  be  later 
hoisted  out  with  a  crane.  A  large  bed  of  black  copper  was  obtained 
before  it  was  possible  for  any  of  the  ore  that  had  been  charged 
after  the  slag-charges,  to  get  down.  The  now  thoroughly  hot  fur- 
nace required  only  1,200  gallons  of  jacket-water  per  hour,  whereas 
in  its  frozen  condition  it  was  using  something  more  than  three 
times  that  amount. 

A  slight  alteration  of  the  ore-mixture  was  found  advisable,  and 
eventually  the  furnace  settled  down  on  to  a  charge  of  300  pounds 
coke,  2,250  pounds  ore,  100  pounds  old  slag,  and  such  foul  slag  as 
was  daily  made  in  the  process.  The  first  tap  of  black  caliper  con- 
tained 32  per  cent.  iron.  After  regular  work  had  become  estab- 
lished the  iron  fell  to  about  4  per  cent. 

In  the  light  of  the  preceding  pages  it  is  not  difficult  to  see  what 
was  occurring  at  the  start.  There  were  several  errors  in  judgment. 
In  the  first  place,  the  furnace  was  started  on  ore  instead  of  on  a 
ferruginous  slag-charge.  This  is  not  absolutely  necessary,  but  it 
makes  things  much  more  comfortable  to  start  with  a  charge  or  two 
of  good,  basic  slag,  and  when  blowing-in  on  new,  untried  ores,  it 


is  doubly  important  to  do  so.  Again,  the  normal,  calculated 
charge  was  used  from  the  outset,  whereas  it  is  always  wise  to  start 
a  siliceous  charge  so  that  the  slag  shall  contain  some  5  per  cent, 
less  silica  than  it  is  eventually  intended  to  keep  it  at.  It  is  easy 
enough  to  make  a  basic  charge  more  siliceous,  but  very  tedious 
and  difficult  to  render  a  chilled,  slow-running,  siliceous  charge 
more  basic. 

The  third  and  greatest  mistake  was  the  use  of  too  much  coke. 
The  charge  only  required  some  12  or  13  per  cent,  of  coke  to  melt  it, 
and  the  extra  12  per  cent  of  the  25  per  cent,  actually  used  could  only 
expend  itself  in  heating  unnecessary  jacket-water  and  in  reducing 
iron  out  of  the  slag.  This  was  exactly  what  occurred.  The  ample- 
feed  pipes  could  barely  supply  the  jackets  with  sufficient  water;, 
the  surplus  heat  ind  strong  blast  forced  the  combustion  to  gradu- 
ally ascend  the  ?,liaft  until,  on  my  arrival,  the  coke  was  burning 
fiercely  at  the  charging  door.  The  metallic  and  most  fusible  por- 
tions of  the  ore  were  liquated  out  in  this  intense  heat;  some  of  the 
iron  was  reduced  to  the  metallic  form,  and  by  tlie  time  the  slow- 
sinking  column  had  reached  the  proper  smelting  zone,  it  was 
merely  a  dry  and  highly  siliceous  skeleton,  with  all  the  coke  burned 
away  in  the  upper  regions  of  the  furnace,  and  ready  to  chill  inta 
a  solid  mass  as  soon  as  touched  by  the  blast.  If  the  furnace  could 
have  been  turned  upside  down,  there  might  have  been  some  chance 
for  it,  as  neavly  all  the  heat  was  in  the  upper  part  of  the  shaft  and 
the  hearth  and  tuyere  zone  were  black  and  cold.  The  metallic 
copper  liquated  from  the  ore  above  trickled  down,  alloying  itself 
with  the  reduced  iron,  and  finally  set  as  a  solid  mass  in  the  chilled 
hearth.  The  only  slag  obtained  came  from  the  slow  liquation 
process  that  was  going  on  several  feet  above  the  tuyeres,  and 
trickled  down  just  back  of  the  breast,  where  it  was  protected  from 
the  blast. 

Evidently,  two  things  had  to  be  done  promptly  if  the  furnace 
was  to  be  saved.     These  were: 

1.  To  get  the  heat  down  from  above,  where  it  was  doing  mis- 
chief, into  the  tuyere  zone,  where  it  belonged. 

2.  To  dissolve  and  remove  the  partly  infusible  skeleton  of  silica 
that  blocked  up  the  tuyeres  and  smelting  zone. 

These  objects  were  accomplished: 

1.  By  greatlv  reducing  the  blast  pressure  to  prevent  unnecessary 
chilling  of  the  half-fused,  cokeless  masses  in  the  hearth,  and  the 
driving  of  the  heat  any  higher  up  the  shaft.     At  the  same   time. 


every  alterDate  tuyere  was  plugged,  to  give  its  cliillecl  nose  a  chance 
to  melt  away. 

2.  By  allowing  the  ore-column  in  the  shaft  to  sink  several  feet, 
and  then  adding  all  at  once  several  full  charges  of  coke  and  ferru- 
ginous slag. 

This  cooled  the  heated  shaft  at  once  (the  uj)per  portion  of  it 
being  of  fire-brick),  and  permanently  suppressed  all  fire  on  top. 
This  suppression  resulted  partly  from  the  light  blast,  partly  from 
the  now  cooled  walls  which  could  not  ignite  the  coke  in  the  upper 
zones,  and  most  of  all  from  supplying  the  new  coke  with  all  the 
work  that  it  could  do,  so  that  it  had  no  heat  left  with  which  to  do 
mischief.  The  basic  slag  that  was  charged  could  not  be  robbed  of 
its  iron,  as  it  smelted  into  a  thin  liquid  long  before  it  could  be 
subjected  to  any  dangerous  reducing  action,  and,  trickling  down 
toward  the  hearth  in  a  multitude  of  thin  streams,  permeated  the 
quartzose  skeleton  opposite  the  tuyeres  in  every  direction.  Taking 
up  additional  silica  with  avidity,  it  acted  very  much  as  a  stream  of 
hot  water  would  act  upon  an  already  fissured  and  rotten  mass  of 
ice.  The  noses  of  the  plugged  tuyeres,  not  having  their  heat  any 
longer  absorbed  by  the  blast,  were  the  first  to  soften  and  disappear, 
and  when  they  became  reasonably  free,  and  the  fresh  coke  had  a 
chance  to  get  down  in  front  of  them,  they  were  put  in  action  and 
their  neighbors  were  plugged  and  given  the  same  chance  to  recu- 

As  soon  as  the  fresh  coke  and  the  ferruginous  slag  got  down  to 
the  tuyere  level  the  action  became  very  rapid,  and  the  great  cliilled 
mass  below,  added  to  the  basic  flux  from  above,  made  a  most  copi- 
ous slag-flow  that  thoroughly  warmed  the  hearth  and  forehearth 
and  heated  up  the  frozen  block  of  copper  that  still  encumbered 
the  crucible,  and  that  was  strongly  alloyed  with  the  iron  that  had 
been  stolen  from  the  mixture  during  the  first  period  of  the  blast. 
This  copper,  owing  to  its  great  heat-conducting  capacity,  remained 
solid  until  its  entire  mass  had  been  brought  to  the  point  of  fusion, 
when  it  melted  all  at  once  and  yielded  the  large  bed  of  ferruginous 
copper  already  mentioned. 

Nothing  now  remained  but  to  regulate  the  mixture  so  as  to  pro- 
duce the  cleanest  and  most  advantageous  slag  possible  under  the 
conditions,  and  then  to  gradually  increase  the  ore-charge  until  the 
coke  was  carrying  every  pound  of  ore  that  it  could  possibly  smelt. 
Then,  and  not  till  then,  was  the  furnace  doing  its  work  properly 


and  employing  its  fuel  in  smelting  ore  instead  of  in  heating  water 
and  reducing  iron. 

If  the  above  description  of  a  most  common,  and  often  disastrous, 
state  of  affairs  seems  a  little  too  minute  for  a  text-book  of  this 
description,  it  must  be  recollected  that  the  illustration  just  given 
deals  with  the  major  portion  of  the  difficulties  encountered  in 
running  a  water-jacket  furnace. 

The  amount  of  water  required  by  a  water-jacket  furnace,  cooled 
from  hearth  to  throat,  depends  so  much  upon  the  local  conditions, 
that  it  is  impossible  to  lay  down  any  fixed  rules  for  its  consump- 
tion. It  will  depend  mainly  upon  the  following  factors,  arranged 
according  to  their  influence: 

Whether  the  ores  smelted  are  of  a  nature  to  form  a  uniform 
protective  coating  upon  the  walls  of  the  shaft  and  one  that  does 
not  grow  so  rapidly  as  to  require  frequent  removal  by  barring. 

Whether  the  coke  used  is  kept  supplied  with  all  the  ore  it  can 
possibly  take  care  of,  so  that  it  may  have  no  energy  left  to  waste 
in  heating  jacket- watei*. 

Whether  the  protective  coating  consists  of  substances  that  are 
good  conductors  of  heat,  or  the  reverse. 

Whether  the  ore  is  granular,  or  contains  too  large  a  proportion 
of  fines,  in  which  latter  case  it  will  be  necessary  to  frequently  allow 
the  ore  column  to  sink  deep  in  the  shaft,  and  thus  expose  the 
jackets  to  powerful,  though  temporary,  heat. 

Upon  the  pressure  of  the  blast  and  the  skill  exercised  in  the  gen- 
eral management  of  the  furnace,  so  that  the  heat  shall  not  keep 
constantly  mounting  toward  the  tunnel-head. 

Upon  the  specific  heat  of  the  slag. 

Under  ordinar)''  conditions  and  proper  management,  the  maxi- 
mum amount  of  feed -water  required  is  shown  in  the  following 
table,  compiled  from  personal  experience,  and  referring  to  furnaces 
when  run  up  to  full  capacity. 

Water  per  hour  while  Water  per  hour  during 

Hearth  Area.  blowing  in  and  out.  normal  running. 

Square  feet.  Galls.  Galls. 

3    900 460 

5    1,200 600 

7    1,450 950 

9.5 2,200 1,100 

12.5 3,000 1,300 

18    4,000 1,.500 

24    5,000 1,800 

30    6,000 2,000 

36    7,000 2,200 


These  figures  refer  to  a  supply  of  fresh  water  ;  but  where  the 
same  water  is  used  over  and  over  again,  about  25  per  cent,  more  is 
required  to  make  up  the  loss  by  evaporation,  etc.,  in  a  3 6 -inch 
furnace  in  the  dry,  hot  climate  of  Arizona. 


The  most  important,  and  apparently  least  understood,  portion 
of  a  copper- matte  blast-furnace  is  the  foreheartli. 

I  have  already  referred  to  the  great  advantage  that  is  gained  by 
allowing  the  molten  products  to  escape  from  tlie  hearth  as  soon  as 
formed,  thus  transferring  the  settling  operation  from  the  inside  to 
the  outside  of  the  furnace.  The  burning-out  of  the  crucible,  the 
formation  of  sows  and  accretions,  and  the  various  difficulties  that 
are  inseparable  from  the  use  of  an  inside  crucible  are  thus  avoided, 
and  even  if  they  are  only  transferred  from  the  interior  to  the  exte- 
rior of  the  furnace,  it  is  an.  enormous  advantage  to  have  them 
where  they  are  distinctly  visible  and  can  be  got  at  and  remedied  at 
once.*  Before  the  days  of  external  forehearths,  more  than  75  per 
cent,  of  the  delays  and  difficulties  encountered  in  running  a  blast- 
furnace were  connected  with  troubles  and  uncertainties  regarding 
the  condition  of  the  hearth  and  crucible,  and  a  furnace  was  often 
run  at  a  loss  for  a  considerable  period,  in  the  hope  that  it  would 
"come  round  all  right  eventually"  and  save  the  cost  and  delay  of 
blowing  out  and  putting  in,  and  drying,  a  new  crucible. 

In  the  modern  practice  these  troubles  are  transferred  to  the  fore- 
hearth,  and  with  proper  arrangements  it  takes  only  30  to  60  min- 
utes to  replace  it  with  a  fresh  one. 

There  are  two  objections  sometimes  urged  against  the  abolition 
of  an  internal  crucible  or  sump,  though  I  have  never  heard  either 
of  them  cited  by  men  who  had  a  varied  experience  in  the  use 
of  suitable  external  settlers.      It  is  sometimes  alleged 

1.  That  there  is  a  loss  of  heat  experienced  by  giving  up  the  in- 
ternal crucible. 

2.  That  the  matte  is  more  perfectly  settled  inside  the  furnace. 
While  it  may  be  possible  to  select  isolated  cases  in  which  either 

*  It  must  be  remembered  that  I  am  referring  entirely  to  American  condi- 
tions, where  100  tons  or  more  of  ore  are  smelted  in  the  furnace  daily,  practically 
without  slag,  and  if  at  all  possible,  without  flux,  and  that  but  two  products 
are  allowable  :  A  matte  of  good  grade  and  often  containing  considerable 
amounts  of  silver  and  gold,  and  a  slag  that  must  be  poor  enough  to  go  over 
the  dump. 


or  both  of  these  objections  might  be  valid,  my  own  experience 
contradicts  them  completely  under  ordinary  conditions.  Where  a 
siliceous  slag  is  made  from  a  hard-smelting  mixture  and  but  a 
small  quantity  of  high-grade  matte  is  produced,  it  is  often  a  ques- 
tion whether  the  inside  crucible  might  not  be  more  economical. 
The  smelting  at  Mansfeld  in  Prussia  is  a  typical  instance  of  this 
kind.  The  ore,  after  being  burued  in  large  heaps  to  remove  the 
bitumen  of  the  shale,  is  smelted  in  large,  high  blast-furnaces,  with 
hot  blast  and  under  conditions  much  resembling  those  present  in 
the  production  of  pig  iron  from  its  ores.  About  17  tons  of  the 
ore  are  concentrated  into  one  ton  of  matte,  having  about  the  fol- 
lowing composition: 

Copper 45  per  cent. 

Sulphur 24 

Iron 20 

Zinc 4.5 

Lead 1 


with  small  amounts  of  manganese,  cobalt,  nickel,  and  silver. 

An  ordinary  slag  from   the  Saengerhausen  smelter  had  the  fol- 
lowing composition,  according  to  Heine: 

Silica 53.83  per  cent. 

Alumina 4.48 

Lime 33.10 

Magnesia 1 .67 

Ferrous  oxide 4.35 

Cuprous  oxide 0.25 

Fluorine 2 . 0» 


With  this  poor  ore,  high  rate  of  concentration,  comparatively 
rich  matte,  and  extraordinarily  siliceous  slag  (probably  the  most 
siliceous  slag  made  regularly  anywhere  in  the  world,  in  blast-fur- 
naces), an  interior  crucible  is,  no  doubt,  essential.  Even  here, 
however,  it  causes  more  or  less  annoyance  and  delays,  as  well  as 
the  production  of  nickeliferous  sows,  which  are  sold  at  the  valua- 
tion of  pig  iron. 

But  for  anything  approaching  ordinary  conditions,  the  two 
objections  cited  are  not  valid  so  far  as  I  am  competent  to  judge. 

The  first  objection,  that  the  use  of  a  forehearth  causes  a  loss  of 
useful  heat,  is  not  difficult  to  meet,  as  it  can   be  almost  anywhere 


decided  by  actual  trial  at  very  slight  exjjense.  I  have  tried  the 
experiment  on  several  occasions  and  with  considerable  care,  and 
have  never  been  able  to  effect  any  saving  in  fuel  by  retaining  the 
matte  in  the  furnace  in  a  crucible,  though  I  have  very  frequently 
witnesseil  a  decided  increase  in  its  consumption  from  irregularities 
brought  about  by  this  practice. 

I  am  uot  aware  that  any  oue  claims  auy  demonstrable  saving  in 
fuel  in  the  actual  smelting  operation  from  the  use  of  the  interior 
crucible.  The  ordinary  statement  is,  that  by  retaining  a  large 
body  of  matte  in  the  furnace,  the  bottom  and  hearth  are  kept 
hot  and  in  good  condition. 

To  this  I  reply,  that  all  I  demand  of  a  bottom  is  to  have  it  fur- 
nish me  a  solid  and  slightly  inclined  surface  on  which  my  molten 
products  can  run  out  of  the  furnace  as  fast  as  they  are  formed, 
and  that  it  is  a  matter  of  entire  indifference  to  me  if  its  tempera- 
ture is  20  degrees  below  zero,  as  I  feel  entirely  confident  that  the 
matte  and  slag  will  simply  chill  enough  to  form  a  non-conducting 
crust  sufficiently  thick  to  prevent  the  cold  bottom  from  stealing 
heat  from  the  constant  and  powerful  stream  of  molten  products. 

A  damp  bottom  is  a  decided  evil,  as  the  escaping  steam  bubbles 
through  the  slag  and,  aside  from  its  cooling  influence,  produces 
serious  trouble  mechanically.  But  a  cold  bottom  need  never  be- 
feared  in  a  fast-running  furnace  with  independent  forehearth. 

As  regards  the  second  objection,  which  has  to  do  with  the  set- 
tling of  the  matte,  I  will  frankly  admit  that  many  forehearths  are 
so  constructed  and  managed  as  to  lose  more  copper  than  would  be 
the  case  with  an  interior  crucible,  always  providing  that  we  could 
guarantee  the  latter  against  irregrlarities.  But  this  loss  in  matte 
comes  almost  entirely  either  from  a  badly-arranged  forehearth,  or 
from  want  of  care  and  skill  in  managing  it.  If  a  furnace  with 
interior  crucible  were  run  with  the  same  carelessness. and  noncha- 
lance that  is  so  often  displayed  in  running  oue  with  forehearth,  the 
question  would  soon  arise,  not  how  to  save  the  matte,  but  how  to 
save  the  furnace.  Because  a  system  is  so  easy  and  comfortable  as 
to  frequently  lead  to  carelessness  and  abuses  is  no  valid  reason  for 
discarding  it  and  refusing  to  profit  by  its  advantages.  The  small 
proportion  of  copper  that  exists  in  the  slag  as  cuprous  oxide  will 
not  be  saved  (by  any  ordinary  means)  either  in  a  crucible  or  in  a 
forehearth,  and  the  matte  globules  themselves  can  be  settled  as 
perfectly  in  the  one  as  in  the  other. 

The  forehearth  in  its  simplest  form  consists  simply  of  some  sort 

284  MODERN'    COPPEK    SMilLTlNG. 

of  vessel  oi*  box  external  to  the  shaft  of  the  furnace,  into  which 
the  molten  products  can  flow,  and  separate  according  to  their  spe- 
cific gravity.  A  simple  settling-pot  is  a  forehearth,  though  a  rude 
and  unsatisfactory  one. 

This  subject  is  so  important  to  the  metallurgist  that  I  shall  de- 
scribe the  main  forms  of  forehearths  in  detail,  premising  that  their 
variety  is  considerable,  and  that  each  form  described  stands  merely 
for  a  general  type  that  has  many  variations.  I  shall  neglect  the 
rudest  and  simplest  forms  of  forehearth,  as  they  are  inefficient  and 
can  also  be  easily  deduced  from  the  more  perfect  ones. 

Perhaps  the  simplest  form  of  efficient  forehearth  is  a  rough'  rect- 
angular box  made  of  four  cast-iron  plates  set  on  edge  on  a  cast-iron 
base-plate,  the  latter  being  mounted  on  wheels,  that  the  whole 
structure  ma)'  be  easily  removed  or  replaced.  The  fastenings  of 
the  plates  should  be  as  simple  as  possible,  being  usually  confined 
to  a  couple  of  rods  that  connect  the  exti'emities  of  the  two  longer 
plates,  the  short  end-plates  being  retained  in  place  by  a  vertical 
ledge  cast  on  the  side-plates. 

There  is  a  removable  cast-iron  slag-spout  at  one  end  and  a  verti- 
cal slot  for  a  tap-hole  at  the  side.  When  the  matte  is  low-grade 
and  plentiful,  and  thus  likely  to  burn  away  the  lining  of  the  fore- 
hearth, the  tap-hole  should  be  situated  near  the  end  of  the  box 
farthest  from  the  furnace.  When  the  matte  is  scanty  or  tolerably 
high-grade,  and  thus  liable  to  chill,  the  tap-hole  should  be  placed 
nearer  the  furnace,  that  it  may  receive  all  the  heat  possible  from 
the  molten  stream  that  is  constantly  entering  the  forehearth. 

The  lining  of  this,  or  any,  forehearth  must  be  suited  to  the  local 
conditions.  Where  the  slag  is  siliceous,  or  the  furnace  is  small  or 
runs  slowly,  or  the  matte  scanty  and  high-grade — 45  per  cent,  or 
over — the  material  for  the  lining  may  be  found  in  the  nearest  bank 
of  clay  or  loam.  This,  mixed  with  chopped  straw  or  horse  niRiiure, 
as  hair  is  mixed  with  mortar  for  plastering  walls,  forms  a  cheap 
and  satisfactory  lining.  Its  sole  duty  is  to  keep  the  molten  con- 
tents from  cracking  or  burning  the  cast-iron  forehearth  plates 
until  they  begin  to  form  a  protective  crust  upon  the  bottom  and 
sides,  that  will  gradually  continue  thickening  until  the  cavity  of 
the  forehearth  becomes  too  small  to  act  as  a  suitable  settlei",  or  it 
becomes  too  difficult  to  drive  a  bar  through  the  tap-hole.  This 
may  occur  in  a  few  days,  weeks,  or  months.  Perhaps  the  average 
life  of  a  good  sized  forehearth  under  favorable  conditions  may  be 
put  at  three  weeks. 


When  the  matte  is  of  lower  gratle  and  abundant,  and  the  slag 
ferruginous,  such  a  lining  would  be  quickly  destroyed,  and  it  be- 
comes necessary  to  use  4^  inches  of  fire-brick,  laying  the  brick  as 
carefully  and  with  as  thin  joints  as  would  be  done  in  the  vvalls 
of  a  furnace.  Even  this  brick  lining  is  occasionally  insufficient, 
and  where  a  very  large  proportion  of  matte  is  formed,  and  espe- 
cially if  the  furnace  is  run  very  hot  in  order  to  store  up  a  large 
charge  of  matte  for  the  Bessemer  converters,  and  at  the  same  time 
keep  this  great  body  of  matte  at  the  high  temperature  required  for 
tapping  into  a  ladle  and  conveying  and  pouring  it  into  the  con- 
verter, it  is  found  necessary  to  greatly  increase  the  size  of  the 
forehearth.  This  is  done  in  tiie  Boston  &  Montana  smelter  at 
Great  Falls,  Montana,  where  charges  of  5  tons  of  molten  matte 
are  required  for  the  converters. 

The  charge  smelted  in  these  blast-furnaces  is  exceedingly  hot 
and  fusible,  consisting  of  50  per  cent,  of  a  mixture  of  first-class 
ore  and  concentrates  assaying  over  20  per  cent,  copper,  36  per  cent. 
converter  slag,  14  per  cent,  limestone,  with  the  addition  of  a  cer- 
tain amount  of  refinery  slag.  The  furnaces  are  36  inches  by  108 
inches  at  the  tuyeres,  and  smelt  about  110  tons  of  the  mixture 
daily.  The  concentration  of  the  ore  is  only  a  little  more  than  2 
into  1,  the  added  slags  enriching  the  matte  considerably,  and  over 
25  tons  of  50  per  cent,  matte  is  produced  daily  at  each  furnace. 

To  enable  the  forehearth  to  store  up  some  5  to  8  tons  of  this 
matte,  it  must  necessarily  be  large;  and  to  prevent  so  much  fiery 
material  from  cutting  out  and  breaking  through  the  lining,  the 
forehearth  has  been  made  shallow  and  has  gradually  been  increased 
in  diameter  (its  form  being  circular),  until  enough  surface  was 
gained  to  chill  the  matte  at  the  circumference  Just  sufficiently  to 
prevent  its  bursting  through  the  lining.  This  is  another  example 
of  the  great  principle  of  automatic  regulation  by  radiation,  that 
has  already  been  prominently  noticed.  At  first  glance  it  might  be 
supposed  that  the  larger  the  forehearth  the  smaller  would  be  its 
radiating  surface  in  proportion  to  its  capacity;  and  this  supposi- 
tion, regarded  simply  as  a  mathematical  proposition,  would  be 
quite  correct.  But  other  factors  modify  the  conditions.  The 
cutting-out  of  a  forehearth  is  effected  largely  by  the  direct  influ- 
ence of  the  constant  white-hot  stream  of  metal  from  the  furnace; 
and  the  farther  the  walls  of  the  forehearth  are  removed  from  this 
center  of  heat,  the  less  will  be  its  infinence  upon  them.  Besido<5, 
in  these  large  forehearths,  the  deptli  of  the  metal  is  very  slight. 


•compared  with  its  extent,  and  tlie  effective  radiating  surface  is 
thus  very  largely  increased. 

These  forehearths  are  now  made  10  feet  in  diameter  over  all, 
the  former  8-foot  ones  having  burst  out  too  frequently.  They 
consist  of  a  cylinder  of  boiler  iron  10  feet  in  diameter  and  without 
top  or  bottom.  The  bottom  is  formed  by  a  course  of  4^  inches  of 
fire-brick  laid  on  a  foundation  of  rammed  clay.  The  lining  con- 
sists of  two  4^-inch  rings  of  fire-brick,  so  that  the  forehearth, 
when  completed,  has  a  clear  diameter  of  nearly  8^  feet.  There 
are  two  tap-hole  slots;  one,  halfway  up  the  side;  the  other  near 
the  bottom.  Of  course  there  is  a  tap-hole  slot  in  the  boiler-iron 
casing  and  a  corresponding  notch  in  the  brick  lining,  but  the  tap- 
hole  proper  is  formed  and  kept  in  condition  by  a  plate  of  copper  as 
high  as  the  forehearth,  2^  inches  thick  at  the  top,  and  tapering  to 
1^  inches  at  the  bottom.  This  slab  of  copper  is  simply  slipped 
•between  the  iron  casing  and  the  brick  lining  and  is  pierced  by  a  1^- 
iuch  circular  hole  that  corresponds  to  the  tap-hole  slot.  Radia- 
tion again  comes  into  play,  and  tbis  uncooled  copper  slab  answers 
its  purpose  satisfactorily  and  keeps  a  free  and  easily  controlled 
tap-hole.  As  a  chill  is  very  apt  to  form  at  the  interior  orifice  of 
the  tap-hole,  especially  as  the  forehearth  gradually  fills  up  with 
accretions,  a  steel  bar  is  kept  constantly  in  the  taphole.  By  driv- 
ing it  slightly  with  a  hammer  from  time  to  time,  its  point  is  kept 
about  even  with  the  inner  surface  of  the  slowly  increasing  chill,  so 
that  the  latter  is  easily  penetrated  by  a  few  solid  blows  when  the 
time  for  tapping  has  arrived. 

The  short  spout  that  conveys  the  molten  products  from  furnace 
to  forehearth  is  of  water-jacketed  boiler  iron,  as  unprotected  metal 
of  any  description  would  be  destroyed  within  an  hour  by  the  pow- 
erful stream  of  matte  and  slag.  The  forehearth  lasts  about  a 
month,  and  handles  some  3,500  tons  of  melted  material  before 
requiring  to  be  replaced. 

It  is  important  with  this,  as  with  every  roofless  forehearth,  to 
get  it  well  protected  at  the  start  with  a  proper  covering  that  shall 
retain  the  heat  and  guard  its  contents  against  too  rapid  chilling. 
This  is  best  effected  as  follows:  A  hot  fire  is  kept  in  the  fore- 
hearth during  the  blowing-in,  and  the  slag-notch  of  the  furnace  is 
not  opened  at  all  until  the  interior  is  filled  pretty  nearly  to  the 
tuyeres  with  molten  products.  The  furnace  should  always  be 
started  on  a  somewhat  basic  and  easilv-fusible  mixture,  and  one 
■that  will  prod  use  a  pretty  large  proportion  of  matte.     Just  before 


the  notch  is  opened,  the  forehearth  shonld  be  scraped  tolerably 
olean  of  ashes  and  cinders,  and  after  the  first  flash  of  slag  has  run 
out  of  the  furnace,  the  surface  of  the  molten  mass  in  the  forehearth 
should  be  liberally  covered  with  light  wood,  the  hot  flame  from 
which  will  prevent  a  stiff  slag-crust  from  forming  on  the  gradually 
rising  bath.  When  the  forehearth  is  full,  it  is  well  to  dam  its 
slag-spout  with  a  lump  of  clay  and  allow  the  bath  to  rise  even 
above  its  surrouuding  walls,  breaking  up  the  slag-crust  all  over 
the  top  with  a  bar,  to  permit  its  free  elevation.  At  length  when 
the  forehearth  is  brimming  full  and  the  center  has  even  risen  two 
or  three  inches  above  the  side  walls,  the  crusted  surface  of  the 
bath  is  evenly  and  thickly  covered  with  a  non-conducting  layer  of 
coke  dust.  This  crusted  roof  of  slag  is  of  the  greatest  importance 
to  the  integrity  of  the  health,  and  should  not  fall  in  when  the 
matte  is  removed  by  tapping.  If  it  should  grow  too  thick  or  too 
thin,  we  only  have  to  apply  our  familiar  principle  of  regulation  by 
radiation  and  add  to,  or  take  away  from,  its  protecting  layer  of 
coke  dust. 

It  is  the  constant  attention  to  just  such  apparently  trifling  de- 
tails as  these  that  enables  some  foremen  to  run  a  furnace  without 
delays  or  difficulties,  while  others  have  frequent  stoppages  and 
constant  trouble  and  hard  work  for  their  men. 

The  next  type  of  forehearth  is  one  rendered  familiar  to  many 
smelters  by  Herreshoff's  patent  water-jacket  furnace.  It  is  a 
much  more  ingenious,  and,  for  certain  conditions,  a  riiuch  more 
perfect  device  than  any  of  those  hitherto  described.  Eun  by  expe- 
rienced furnace-men  and  on  proper  ores,  it  enables  a  furnace  to  run 
as  nearly  absolutely  without  stoppages  or  without  producing  foul 
stag,  as  is  well  possible. 

To  understand  IlerreshofE's  furnace  and  forehearth,  it  will  be 
necessary  to  turn  to  Figs.  27  and  28  on  page  2G5. 

The  furnace  here  shown  is  rectangular  in  shape,  with  corners 
rounded,  and  the  lines  between  the  corners  slightly  curved  or  of 
convex  shape.  The  height  is  10  feet,  width  .3  feet  7  inches  at 
the  bottom,  and  4  feet  7  inches  at  the  top,  by  6  feet  4  inches  length 
at  the  bottom,  and  7  feet  4  inches  at  the  top.  The  water-jacket 
is  exceptionally  narrow,  having  a  water-space  of  only  2  inches. 

Referring  to  the  cuts,  A  is  the  body  of  the  furnace;  B  a  ring  2 
by  2  inches,  to  which  the  plates  of  the  water-jacket  are  riveted. 
At  the  top  C,  the  outer  plate  is  flanged  2  inches,  and  the  inner 
plate  4  inches,  and  the  flanges  then  riveted.     The  bottom  of  the 


furnace  E  is  a  dishetl  cast-iron  plate  1^  inches  thick,  fastened  H 
the  ring  B  by  tap-bolts.  This  permits  the  dropping  of  the  bottom 
if  required.  The  legs  F  are  bolted  to  the  ring  B  on  the  outside  of 
the  furnace,  thus  not  interfering  with  the  dropping  of  the  bottom. 
The  hole  G  is  the  outlet  of  the  furnace  for  both  slag  and  matte. 
It  is  0  inches  high  and  T  inches  wide  and  made  by  riveting  the 
wrought-iron  frame  H  into  the  shell  of  the  furnace.  The  furnace 
is  blown  by  13  tuyeres,  five  on  each  side  and  three  on  the  back. 
They  are  placed  26  inches  above  the  bottom  plate,  and  are  2  inches 
in  diameter. 

The  construction  of  the  furnace  proper  is  practically  identical 
with  that  of  a  former  round  furnace,  but  the  forehearth  is  consid- 
erably changed.  In  the  round  furnace  the  forehearth  was  floored 
with  a  layer  of  slag-wool  and  brick  as  described.  A  brick  lining 
was  also  used.  The  bottom  of  the  brick  lining  was  some  12  inciies 
below  the  outlet  from  the  jacket.  Experience  proved  that  this 
bottom  invariably  chilled  to  a  level  with  the  bottom  of  the  opening 
to  the  furnace.  The  cutting  of  the  brick  lining  at  a  higher  level 
also  gave  occasional  trouble.  Both  these  faults  are  avoided  in  the 
present  construction.  The  former,  by  raising  the  forehearth  on 
high  wheels  N,  and  making  the  floor  of  the  bottom  lining  within 
2  inches  of  a  level  with  the  bottom  of  the  inlet  L  The  latter,  by 
entirely  casting  aside  fire-brick  lining  and  depending  on  the  circu- 
lar cast-iron  water-jacket  K.  The  tap-hole  R  in  the  shaft  of  the 
furnace  is  used  only  when  blowing  out  to  tap  the  furnace  clean,  or 
sometimes,  for  such  small  quantities  of  black  copper  as  may  be 
accidentally  made.  In  the  forehearth,  the  tap-hole  0  is  the  one 
commonly  in  use.  It  is  made  of  copper,  bolted  to  the  iron  body 
of  the  forehearth  and  is  water-jacketed  similarly  to  the  "  Liilirmann" 
slag  tuyere  of  iron  furnaces.  The  manner  of  operating  it  is  also 
similar.  M  is  the  slag-spout;  W,  a  brick-lined,  dish-shaped  mova- 
ble iron  cover  of  the  forehearth.  When  smelting,  the  well  or 
forehearth  is  wheeled  up  against  the  furnace,  as  shown  in  the  cut, 
and  a  very  small  amount  of  wet  fire  clay  is  placed  on  the  iron  faces 
surrounding  the  holes  G  and  L,  in  order  to  make  a  tight  joint 
between  them. 

In  practical  operation,  after  the  furnace  has  been  properly 
charged,  the  blast  is  let  on.  The  first  cinder  collects  in  the  bot- 
tom of  the  furnace  shaft  proper,  and  accumulates  until  it  reaches 
the  holes  G  and  L.  It  then  overflows  rapidly  into  the  forehearth, 
carrying  matte  with  it.     In  a  short  time,  the  level  of  the  nioltea 


material  rises  above  the  top  of  tiie  hole  L,  and  from  that  time  on- 
ward the  blast  iu  the  furnace  can  no  longer  blow  out  through  L,  and 
is  completely  trapped.  Owing  to  the  pressure  of  blast,  the  level 
of  molten  matte  and  slag  in  tlie  forehearth  is  several  inches  above 
that  in  the  furnace  proper.  Eventually  the  slag-lip  M  is  reached 
by  the  cinder,  which  then  overflows  quietly.  Matte  is  tapped 
periodically  from  the  tapping-notch  0  without  stopping  the  fur- 
nace. Matte  is  never  allowed  to  accumulate  until  it  Overflows  at 
the  slag-lip,  the  practice  being  to  tap  at  stated  intervals.  The 
notch  0  is  opened  by  a  small  steel  bar,  and  pure  matte,  to  the 
amount  of  about  1,000  pounds,  is  allowed  to  run  off.  During  this 
operation,  the  level  of  the  molten  slag  m  the  forehearth  falls,  but 
not  sufficiently  to  admit  of  blast  escaping  through  L. 

By  the  simple  insertion  of  a  small  clay  stopper,  the  matte  is 
stopped  before  cinder  appears,  thus  avoiding  all  cinder  picking. 
The  whole  process  only  occupies  a  few  minutes,  and  is  so  perfect 
that  for  months  a  miss  in  tapping  or  closing  up  is  not  made.  The 
large  amount  of  molten  slag  and  metal  in  the  forehearth  greatly 
facilitates  a  clean  separation. 

While  using  this  forehearth  in  cold  climates  I  have  been  trou- 
bled with  its  frequent  cracking.  On  this  account,  I  have  designed 
a  wrought-iron  one  of  similar  pattern,  but  with  a  separate,  water- 
cooled  slag-spout.  This  has  been  found  entirely  satisfactory  and 
durable,  and  in  spite  of  its  greater  first  cost,  is  the  more  econom- 
ical in  the  end.     This  is  shown  in  Figs.  35,  36,  37  and  38. 

Owing  to  the  narrow  water-cooled  passage  between  furnace  and 
forehearth,  the  Herreshofl:  furnace  requires  careful  management  to 
prevent  the  '*  sticking-up"  of  this  notch.  If  only  temporarily  blocked 
by  a  fragment  of  coke  or  quartz,  it  can  usually  be  cleared  by  prob- 
ing it  with  a  long  rod  of  ^-inch  or  f-inch  iron  passed  through  the 
slag-spout  of  the  forehearth.  In  ordinary  cases  there  is  no  occa- 
sion for  emptying  the  forehearth  for  this  purpose,  but  when  the 
delay  promises  to  exceed  a  very  few  minutes  it  is  safer  to  tap  the 
foreheartb  dry  and  even  to  empty  the  furnace  through  the  little 
notch  provided  therefor. 

The  furnace  can  be  blown  in  or  out  with  great  ease  and  little 
loss  of  time  or  coke,  or  it  can  be  run  simply  on  the  day  shift,  shut- 
ting down  nights  without  blowing  out.  To  do  this  neatly,  it  is 
best  to  reserve  all  the  foul  slag  made  through  the  day,  and  after 
allowing  the  charge  to  sink  tolerably  low  in  the  shaft,  to  charge 
the  slag  and  a  blank  charge  of  coke,  with  a  little  fine  coke  on  top, 

Fig.  35. 


SiPE  ELEV/7r/eW 

Fig.  37. 

FRONr  ei£v//r/o/v 

Fig.  88. 


ami  then,  after  tapping  the  forehearth  and  furnace  completely 
dry,  to  plug  all  the  tuyeres  and  every  other  opening  where  air 
could  penetrate  below.  In  the  morning,  the  blast  is  put  on  at 
once,  and  the  furnace  shaft  rapidly  filled  with  the  regular  charge. 
Slag  will  commence  running  almost  immediately  and  the  smelting 
may,  practically  speaking,  be  taken  up  where  it  was  left  the  night 
before,  except  that  the  first  matte  produced  will  be  a  little  richer 
than  usual  on  account  of  the  slow  roasting  of  the  ore  in  the  furnace 
during  the  night. 

The  Herreshoti  type  of  forehearth  is  not  suited  to  irregular  run- 
ning and  frequent  stoppages,  nor  to  sudden  changes  in  the  ore 
mixture.  Nor  can  it  conveniently  produce  so  siliceous  a  slag  or  so 
high  grade  a  matte  as  a  non-cooled  hearth  and  one  that  permits 
easier  access  to  the  breast  of  the  furnace.  And  above  all,  it  cannot 
be  mauaged  by  inexperienced  men,  as  has  been  proved  more  than 
once  by  its  rejection  at  points  where  it  was  actually  the  ideal  appa- 
ratus for  the  circumstances.  But  for  the  steady  smelting  of  uni- 
form copper  ores,  producing  a  moderately  free  slag  and  a  matte 
between  20  per  cent,  and  50  per  cent.,  it  can  be  run  a  greater 
number  of  hours  in  the  year  and  with  fewer  repairs  and  less  foul 
slag,  than  any  other  furnace  with  which  I  am  acquainted.  It  also 
produces  its  matte  absolutely  free  from  slag  or  other  impurities. 

The  separate  tap-hole  casting  belonging  to  the  Herreshofl  fore- 
kearth  is  a  circular,  water-cooled  bronze  block.  It  is  sometimes 
difficult  to  obtain  these  bronze  castings  perfectly  free  from  flaws 
or  blowholes,  but  the  material  is  malleable  enough  to  stand  a  good 
deal  of  caulking,  and  the  life  of  even  a  bad  tap-hole  casting  may  be 
greatly  prolonged  by  running  its  feed-water  as  hot  as  practicable. 
During  the  moment  that  matte  is  being  tapped,  a  little  more  feed- 
water  should  be  turned  on,  that  steam  may  not  be  generated  in 
the  water  space.  As  this  tap-hole  is  always  plugged  as  soon  as  a 
potful  of  metal  lias  been  run  oti,  the  furnace-man  has  to  plug  it 
against  a  tolerably  strong,  though  small.  Jet  of  metal  that  is  forced 
out  both  by  the  pressure  of  the  blast  and  the  column  of  metal  in 
the  forehearth.  Any  dampness  about  the  clay  plug  or  on  the  tap- 
hole  casting  is  likely  to  cause  a  shower  of  the  molten  matte  to 
blow  back  in  his  face,  and  as  he  naturally  turns  his  head  and 
shrinks  from  the  bombardment,  he  is  likely  to  miss  his  shot. 
Thus  men  are  frequently  burned,  and  matte  is  spilled  about  the 
floor.  This  little  annoyance  is  avoided  by  a  movable,  swinging 
screen  of  s>heet  iron,  having  a  vertical  slot  through  it  for  the  plug- 



ging-pole,  and  a  pane  of  glass  through  which  the  furuace-man  can 
see  what  he  is  doing.  The  main  blast-pipe  is  also  provided  with  a 
weighted  clapper-valve,  and,  at  the  moment  of  plugging,  the  assist- 
ant pulls  a  wire  which  raises  the  valve  and  allows  the  blast  to 
jscape  into  the  air  for  an  instant.  This  relieves  the  pressure  and 
makes  the  plugging  much  easier.  Under  ordinary  circumstances, 
nothing  larger  than  a  cariienter's  hammer  is  required  to  drive  the 
tapping-bar  through  the  clay  plug.  The  1^-iuch  tap-hole  in  the 
bronze  casting  is  permanently  plugged  with  clay,  and  through  the 
center  of  this  clay  stopper,  a  one-half  inch  hole  is  left.  This  is 
the  tap-hole  proper,  and  is  plugged  with  a  fragment  of  plastic, 
slightly  dried  clay  no  larger  than  a  cork.  Outside  of  this  minute 
5>lug  a  larger  plug  is  forced  in,  and  when  the  furnace  is  to  be  tap- 
ped, the  attendant  removes  the  outside  clay  with  a  small  instru- 
ment like  the  blade  of  a  pocket-knife.  When  the  protected  inner 
plug  is  reached,  a  light  tap  or  two  that  might  almost  be  given  with 
the  ball  of  the  hand,  is  sufficient  to  drive  the  half-inch  steel 
tapping-bar  through  the  thin  obstruction. 

The  following  table  gives  a  week's  run  of  one  of  these  furnaces, 
taken  from  the  daily  sheet  of  returns.  It  illustrates  the  steady 
running  and  the  small  amount  of  slag  formed.  As  I  have  selected 
a  week  when  there  was  no  changing  of  forehearth,  washing  out  of 
jacket,  or  important  repairs  of  any  description,  the  record  is  some- 
what better  than  the  average  for  the  year  would  be. 

During  the  week  the  furnace  averaged  110  tons  ore  per  24  hours, 
making  in  the  same  time  about  15  tons  of  40  per  cent,  matte. 




Ore  Smelted. 

















....  771 

Foul  Slag.    Blast  off  Furnace 
Pounds.  Minutes. 







Cause  of  Delay. 
Clearing  slag-bole. 
Patching  forehearth. 

Engine  repairs. 
Ore  train  late. 

Slag  in  tuyeres. 

This  makes  the  delays  amount  to  0.84  per  cent,  of  the  total 
time,  and  the  production  of  foul  slag  equals  about  one-third  of  one 
per  cent,  of  the  ore  smelted.     It  need   hardly  be  said  that  the 



ores  were  exceediugly  uniform  and  favorable,  the  plant  excellent, 
and  the  furnace-men  thoroughly  experienced  and  interested  in  the 

The  Orford  siphon-tap*  forehearth  is  an  outside  settling  device 
so  arranged  that  the  matte  and  slag  are  discharged  from  it  in  sep- 
arate and  continons  streams.     See  Figs.  39  and  -iO. 

It  consists  of  a  rectangular  box,  some  5  feet  by  5  feet  6  inches, 

Figs.  39  and  40. 

formed  of  cast-iron  plates  strongly  bolted  together  at  the  corners, 
and  lined  with  a  brick  wall  44^  inches  or  9  inches  thick,  according  to 
the  quality  of  the  product.  It  is  fastened  firmly  to  the  front  of 
the  furnace,  just  at  the  slag-run  in  tlie  center  panel,  the  lower 
middle  portion  of  the  anterior  front  wall  of  that  structure  forming 

•  This  is  an  entire  misnomer,  as  the  apparatus  here  referred  to,  a.«  used  for 
the  continuous  discbarge  of  tlie  metallic  product,  has  nothing  about  it  pertain 
ing  to  tlie  principles  of  the  siphon. 


its  posterior  boundary.  It  is  divided  longitudinally  by  a  '.)-iucli 
wall  of  fire-briok  into  a  greater  and  lesser  portion,  the  area  of  the 
two  compartments  being  about  as  5  to  2,  and  the  direction  of  the 
division  wall  bring  parallel  to  the  short  axis  of  the  furnace 

The  entire  molten  contents  of  the  furnace  discharge  through  a 
2-inch  by  4-iuch  opening  (the  slag-run)  in  the  middle  panel  (the 
breast)  into  the  larger  of  these  two  compartments,  which  is  pro- 
vided with  a  slag-spout,  bolted  to  the  upper  edge  of  the  front  plate, 
while  it  communicates  with  the  smaller  compartment  by  means  of 
a  3-inch  by  8-inch  vertical  slot  through  the  9-inch  division  wall, 
about  midway  of  its  length  and  on  a  level  with  the  floor  of  the 
forehearth.  This  smaller  compartment  also  has  a  spout  about  2 
inches  below  the  level  of  the  spout  belonging  to  the  larger  division, 
and  on  the  outer  side,  instead  of  the  end  wall,  for  the  sake  of 

A  thorough  understanding  of  this  very  simple  and  inexpensive 
contrivance  will  render  it  very  easy  to  appreciate  its  management. 

When  the  breast-hole  is  opened,  and  slag  and  metal  first  begin 
to  flow,  the  larger  compartment  is  soon  filled,  as  the  only  means  of 
communication  between  the  two  divisions  of  the  forehearth  is  the 
closed  slot  in  the  lower  part  of  the  9-inch  division  wall. 

The  molten  products  separate  according  to  the  law  of  gravity, 
and  slag  is  allowed  to  flow  through  the  spout  of  the  large  compart- 
ment until  the  drops  of  metal  appearing  show  that  it  is  filled  with 
the  more  valuable  product.  The  channel  of  communication  is  now 
opened  by  means  of  a  crooked  tapping-bar,  and  the  metal  flows 
rapidly  through  the  same  into  the  smaller  compartment,  until  an 
equilibrium  is  established,  and  both  divisions  of  the  forehearth  are 
partially  filled  with  the  matte,  the  communicating  channel  being 
far  below  the  surface  of  the  same,  and  consequently  so  situated 
that  slag  can  never  reach  it  unless  it  should  sink  below  the  metal, 
which  is  obviously  impossible. 

As  the  furnace  constantly  discharges  its  stream  into  the  larger 
compartment,  the  forehearth  is  soon  filled  again,  the  metal  sinking 
to  the  bottom  and  standing  at  the  same  level  in  both  divisions, 
while  the  slag  simply  flows  over  the  surface  of  the  matte  in  the 
larger  compartment. 

As  soon  as  the  matte  reaches  the  level  of  the  spout  attached  to 
the  small  compartment,  it  begins  to  flow  into  a  pot  placed  to  receive 
it,  and  by  judicious  manipulation,  and  if  a  sufficient  proportion  of 


Fu;.  42. — Section  through  Vertical. 


matte   is  produced  from   the  charge,  a  constant  stream   of  each 
product  may  be  kept  running  without  difficulty. 

TJie  management  of  this  siphon-tap  requires  considerable  expe- 
rience, as  the  matte  stops  occasionally  without  apparent  cause, 
and  requires  a  certain  amount  of  manipulation  and  coaxing  to 
keep  running  freely.  This  is  accomplished  by  slightly  damming 
up  the  slag-spout,  which  soon  forces  an  excess  of  matte  into  the 
smaller  compartment,  or  by  cheariug  out  the  communicating  orifice 
by  means  of  a  heated  bar  bent  to  the  required  curve. 

With  matte  of  50  per  cent,  or  over,  the  principal  difficulty  is 
found  in  the  gradual  filling  up  of  the  forehearth  by  chilling,  while 
the  matte  containing  20  per  cent,  or  less  of  copper,  and  produced 
in-  large  quantities,  has  directly  the  opposite  effect,  thinning  the 
fire-lining  until  the  plates  are  endangered,  and  cutting  away  the 
division  wall  until  the  two  compartments  are  virtually  thrown 
into  one. 

But  even  under  these  circumstances,  and  as  long  as  a  vestige  of 
the  center  wall  remains,  the  separation  of  the  matte  and  slag  con- 
tinues to  be  perfect,  and  by  judicious  repairing  and  nursing,  a 
forehearth  apparently  in  the  last  stage  of  ruin  may  yet  do  good 
service  for  many  days. 

An  opening  through  the  division  wall  18  inches  high  by  24 
inches  wide,  and  actually  involving  two-thirds  of  the  separating 
brick-work,  is  not  incompatible  with  a  perfect  separation. 

The  larger  compartment  is  provided  with  a  tap-hole  at  its  lowest 
boundary,  and  on  the  side  opposite  the  matte  division,  and  a  large 
quantity  of  sand  should  always  be  at  hand  ready  to  make  up  into 
rough  molds  in  case  of  any  sudden  necepsity  for  tapping. 

Mafhew.son\s  device  (see  Figs.  41  and  42)  for  separating  matte 
and  slag  has  usually  been  applied  to  lead-silver  blast-furnaces 
where  the  matte  is  of  very  secondary  importance.  It  may,  however, 
prove  useful  to  the  copper  smelter  where  exceptional  circumstances 
demand  the  employment  of  an  interior  crucible,  and  where  the 
amount  of  matte  produced  is  very  small  and  used  primarily  as  a 
collector  of  the  precious  metals. 

I  have  seen  this  apparatus  doing  most  excellent  work  in  Pueblo 
and  elsewhere.  The  illustrations  are  taken  from  a  paper  by  B. 
Sad  tier,  in  TJie  Scientific  Qitarfe7-ly,  for  June,  1893. 

The  matte  is  tapped  from  the  lowest  hole  in  the  section,  and 
should  be  free  from  slag.  There  is  a  cleaning  hole  above  this, 
which  is  ordinarily  closed.     The  slag  flows  out   under  a  water- 


jacketed  diaphragm,  and  through  a  spout  which  starts  at  nearly 
the  level  of  the  tuyeres. 

Reverter atory  Forehearths* — By  a  reverberatory  forehearth,  I 
mean  an  independent  settling  reservoir  iuto  which  is  discharged 
the  molten  material  from  the  blast  furnace,  and  which  is  heated 
from  an  independent  source.  This  far,  it  has  been  found  conven- 
ient to  build  this  settler  in  the  shape  of  a  small  reverberatory 

To  save  time  and  repetitious,  it  will  be  advantageous  to  consider 
the  reverberatory  forehearth  from  the  point  of  view  of  both  the 
blast-furnace,  and  the  converter  departments. 

Every  metallurgist  who  is  in  the  habit  of  running  copper  blast- 
furnaces at  a  rapid  rate  on  tolerably  uniform  ores  is  aware  that 
the  larger  portion  of  his  delays,  and  outside  costs  and  losses,  are 
connected  with  the  settling  of  the  matte  from  the  slag. 

If  he  uses  an  inside  crucible,  he  is  likely  to  experience  the  train 
of  evils  already  considered. 

If,  according  to  ordinary  American  practice,  he  employs  an  in- 
dependent forehearth,  he  betters  his  condition  decidedly,  but  is 
still  frequently  annoyed  by  the  burning-out  or  chilling-up  of  the 
forehearth,  the  carelessness  of  the  workmen  in  allowing  matte  to 
run  over  with  the  slag,  and  various  other  evils.  These  irregulari- 
ties come  largely  from  faults  on  the  part  of  the  furnace-men,  and 
occur  so  much  more  frequently  during  the  night  shift,  that  I  have 
been  in  the  habit  of  saying,  for  instance,  that  under  certain  speci- 
fied conditions,  my  forehearths  would  last  for  20  night  shifts  or 
40  day  shifts. 

The  Bessemer-converter  foreman  may  properly  demand  that  his 
molten  charges  of  matte  shall  be  prepared  for  him: 

[a)  At  the  moment  he  is  ready  for  them;  and  he  may  often 
need  a  double  charge  or  charges  for  two  or  more  converters  in 
rapid  succession. 

*  While  I  have  long  believed  in  reverberatory  forehearths,  and  have  lately- 
bad  opportunities  to  satisfy  myself  of  their  economy  and  effectiveness,  1  find 
that  Dr.  lies,  of  Denver,  has  pursued  the  same  subject  with  much  more  care 
and  thoroughness  than  I  have  ever  devoted  to  it,  and  has,  indeed,  patented  a 
device  of  the  kind.  My  present  object  in  discussing  this  form  of  forehearth  is 
simply  to  point  out  its  possible  value  to  copper  metallurgists,  and  not  to  make 
any  claims  of  either  precedence  or  originality  in  the  matter.  At  the  Messrs. 
Elliott's  Company's  works  in  Wales,  Christopher  James  is  using  reverberatory 
forehearths  with  advantage. 


{b)  So  that  the  matte  has  a  sufficiently  high  temperature  to 
warm  up  a  couverter  that  has  become  too  cool  in  the  preceding 

(c)  That  the  charging  shall  be  accomplished  quickly,  else  both 
the  converter  and  the  slowly  trickling  new  charge  may  become 
unduly  cold. 

(f/)  He  may  desire  to  suspend  using  matte  for  a  considerable 
time,  and  then  require  several  charges  almost  simultaneously. 

(e)  Athough  it  is  a  luxury  he  has  never  been  much  accustomed 
to,  it  would  be  highly  advantageous  if  he  could  order  his  matte 
richer  or  poorer  (within  a  10  per  cent,  limit),  according  to  the 
condition  of  the  lining  of  his  converters. 

These  demands,  and  various  other  causes,  make  it  practically 
impossible  to  attempt  to  tap  the  matte  directly  from  the  ore  blast- 
furnaces into  the  converters.     As  will  be  readily  seen: 

(a)  One  cannot  always  arrange  to  have  a  forehearth  full  of  matte 
at  just  the  moment  that  a  converter  requires  a  charge. 

(b)  It  is  impossible  to  change  the  ore  mixture  in  the  furnaces 
without  disturbing  the  matte-ratio  between  furnaces  and  con- 

(c)  The  best  managed  blast-furnaces  have  their  periods  of  depres- 
sion and  of  exhilaration,  which  decidedly  modify  the  amount  of 
matte  that  they  produce. 

(d)  The  blast-furnaces  must  be  run  at  a  temperature  considera- 
bly above  that  actually  required  to  fuse  the  ore,  in  order  to  keep 
the  matte  in  the  forehearth  sufficiently  hot  for  the  converters. 
The  consequent  waste  of  fuel  is  a  steady  and  considerable 

(e)  If  there  is  any  delay  at  the  converters,  it  reacts  directly  upon 
the  blast-furnaces,  as  they  cannot  dispose  of  their  matte  except  by 
tapping  it  to  one  side  and  remelting  it  later. 

(/)  It  diverts  the  blast-furnace  foreman  from  his  proper  aim; 
which  is  to  smelt  as  much  ore  as  possible  with  the  least  fuel  and 
the  smallest  losses.  He  has  to  constantly  consider  the  needs  of 
the  converters,  and  unduly  push,  or  hold  back,  his  furnaces,  which 
circumstance  is  ruinous  to  economical  smelting,  and  also  affords 
him  an  admirable  and  unanswerable  excuse  for  any  description  of 
accident  or  bad  work.  It  also  destroys  the  spirit  of  rivalry  and 
ambition  which  is  so  important  a  factor  in  large  works. 

(g)  It  causes  endless  complications  and  disputes  between  furnace 
and  converter  departments,  as  each  is  naturally  looking  out  for  its 


own  interests  with  a  total  disregard  for  its  neighbor's  convenience 
or  economy. 

In  view  of  all  these  drawbacks,  experience  has  shown  it  to  be 
more  advantageous  to  go  to  the  considerable  delay  and  expense  of 
breaising,  transporting,  and  re-smelting  the  blast-furnace  matte  in 
a  separate  cupola  that  can  devote  its  entire  attention  and  interests 
to  the  needs  of  the  converters.  This  naturally  entails  a  heavy  ad- 
ditional expense,  amounting,  in  Montana,  to  something  like 
*"2.50  per  ton  of  matte,  besides  requiring  an  investment  for  plant 
of  at  least  §300  for  each  ton  of  matte  melted  per  "24  hours.* 

It  is  also  a  highly  unreasonable  and  aggravating  practice  to 
deliberately  cool  matte  that  is  all  ready  for  the  converters,  and  to 
resmelt  it  again  with  the  consequent  loss  of  labor,  fuel,  time,  and 

Long  before  the  days  of  bessemerizing  copper,  it  had  occurred 
to  certain  metallurgists  that  the  separation  of  the  matte  and  slag 
might  be  facilitated  by  heating  these  products  in  a  separate  reser- 
voir outside  of  the  blast-furnace,  and  by  means  of  an  independent 
fire.  Since  the  almost  universal  adoption  of  independent  fore- 
hearths,  and  especially  since  the  development  of  the  converter 
practice,  the  need  for  such  an  independently  heated  settling- 
reservoir  has  greatly  increased,  as  may  easily  be  gathered  from  the 
brief  explanations  just  given.  Without  attempting  to  speak  of 
tlie  origin,  history,  or  development  of  this  idea,  I  will  state  my 
own  views  as  to  what  seems  to  me  the  most  convenient  form  of 
device  for  this  purpose,  and  the  chief  advantages  that  may  accrue 
from  its  use. 

As  the  management  of  the  reverberatory  forehearth  must  be 
studied  in  conjunction  with  the  running  of  the  blast-furnace  and 
converter-departments,  so  do  its  construction,  maintenance,  and 
repairs  belong  to  the  reverberatory  section.  It  is  simply  a  small 
reverberatory  placed  near  the  blast-furnace,  and  having  the  position 
of  its  fire-box  changed  to  the  side,  instead  of  the  end  of  the  hearth; 
this  modification  is  of  coui"se  not  essential,  but  usually  seems  more 
convenient.  The  blast  furnace  discharges  its  melted  products  into 
the  hearth  of  the  reverberatory  through  an  opening  in  the  rear 
wall  of  the  latter,  and  the  clean  slag  flows  ofiE  continuously  at  the 

*  The  Boston  and  Montana  Company,  at  Great  Falls,  tap  tbeir  matte  into  the 
converters,  via  an  electrically-movei  ladle,  direct  from  the  blast-furnaces  and 
reverberatories.  But  few  concerns  have  either  the  rich  ores  or  the  large  capital 
necessary  to  arrange  a  plant  satisfactorily  on  these  lines. 


front,  or  skimming-door  end.  The  matte  is  tapped  into  the  con- 
verters, either  direct,  or  through  the  intervention  of  a  ladle. 

Very  little  need  be  said  about  the  construction  of  this  forehearth. 
The  ordinary  reverberatory  furnace  offers  a  perfect  model,  and  the 
only  changes  required  are  those  necessary  to  adapt  it  to  its  peculiar 

In  planning  its  position  in  regard  to  the  blast-furnace,  the  fol- 
lowing ^joints  should  be  borne  in  mind  : 

1.  To  have  it  convenient  for  the  removal  of  the  slag. 

2.  To  arrange  it  so  that  the  matte  can  be  tapped  direct  into  any 
one  of  the  converters  (unless  a  ladle  is  used),  and  also,  to  have 
ample  room  to  tap  a  very  large  charge  of  matte  into  sand  beds  at 
one  side,  and  plenty  of  space  to  store  50  to  100  tons  of  matte  in 
pigs.  ^ 

3.  To  so  j)lace  the  forehearth  that  the  breast  of  the  blast-furnace 
can  be  easily  and  freely  reached  with  tools. 

4.  To  so  plan  it  that  the  supply  ot  co:il  for  the  forehearth  can 
be  ec'onomically  and  conveniently  delivered  and  stowed. 

5.  To  so  plan  the  reverberatory  stack,  or  down-take,  that  it 
may  not  be  in  the  way  and  will  not  involve  a  too  expensive  con- 
struction of  flues. 

This  little  reverberatory  should  be  constructed  with  a  fire-box 
that  can  be  comj)letely  closed,  as  in  the  long  calciner  shown  in 
Fig.    22.     This  effects  a  considerable  saving  in  fuel. 

After  the  experience  of  Griffiths  &  James  in  Wales,  and  similar 
practice  at  Mausfeld  and  elsewhere,  no  one  should  think  of  using 
a  sand,  or  quartz  iiearth  in  such  a  settling  reverberatory.  A 
slightly  concave  bottom  of  ordinary  Stourbridge  brick  has  already 
lasted  two  years  in  such  a  forehearth,  running  on  very  foul  and 
leady  mattes,  and  shows  no  signs  of  wear. 

The  only  portion  of  the  reverbei-atory  that  may  require  occa- 
sional looking  after  is  where  the  surface  of  the  slag  touches  the 
fettling.  At  this  point  it  is  liable  to  cut  a  groove  all  around  the 
hearth,  owing  to  its  solvent  action  on  the  silica  of  the  lining. 
Consequently,  the  hearth  may  require  a  little  claying  once  in  from 
three  to  ten  days.  By  surrounding  the  hearth  at  this  level  with  a 
H-inch  pipe,  using  about  200  gallons  of  water  per  hour,  I  have 
almost  entirely  prevented  the  destruction  of  the  lining.  A  suffi- 
cient crust  of  accretions  is  formed  outside  of  the  pipe  to  protect 
the  walls  very  completely. 

The  size  and  depth  of  the  hearth  must  depend  upon  the  weiglit 


of  the  charge  required  for  the  converter,  aud  the  number  of  these 
vessels.  lu  auy  case,  a  large  body  of  matte  kept  constantly  in  the 
hearth  maintains  the  latter  at  a  uniform  heat  and  acts  as  an  excellent 
balance  wheel  for  the  entire  process.  Fifteen  to  thirty  tons  of 
matte  are  none  too  much,  as  there  is  no  difficulty  in  constructing 
a  hearth  that  can  stand  double  that  amount,  providing  it  is  prop- 
erly built  and  ironed.  The  fire-box  may  be  quite  small,  say  30  by 
42  inches,  as  the  amount  of  heat  that  is  required  in  addition  to 
that  already  provided  by  the  molten  products  of  the  blast-furnace, 
is  very  small.  When  slag  is  nearly  hot  enough,  a  rise  of  temper- 
ature of  a  very  few  degrees  makes  an  enormous  difference  in  its 
physical  condition,  and  may  change  it  from  a  cold,  red,  sluggish, 
semi-viscid  substance  to  a  white,  smoking,  oily  liquid,  as  thin  as 
milk.  Besides,  the  conditions  here  are  totally  different  from  those 
that  prevail  in  a  reverberatory  smelting  furnace.  In  the  latter, 
the  greater  proportion  of  the  fuel  is  consumed,  not  in  actually 
melting  the  ore,  but  in 

{a)  Restoring  the  furnace  to  its  normal  temperature  after  it  has 
been  cooled  off  by  skimming,  tapping,  fettling,  charging,  etc. 

{b)  Penetrating  the  feebly-conducting  materials  of  the  charge 
to  reach  the  deeper  layers. 

((■)  Raising  half-molten  masses  from  the  bottom,  where  they 
often  stick  for  a  long  time  after  the  rest  of  the  charge  is  ready  to 

All  these  factors  are  absent  in  the  reverberatory  forehearth.  It 
is  never  cooled  off  by  charging,  skimming,  claying,  or  opening 
doors,  excepting  on  the  rare  occasions  when  the  hearth  requires 
ten  minntes'  repairing.  There  is  no  non-conducting  heap  of  ore 
to  be  penetrated  by  the  heat,  nor  any  half-fused  masses  sticking  to 
the  bottom,  and  there  is  a  constant  stream  of  white-hot  matte  and 
slag  entering  the  forehearth. 

It  is  quite  practicable  to  make  one  forehearth  serve  for  two  or 
more  blast-furnaces. 

The  advantages  offered  by  some  such  form  of  reverberatory 
forehearth  have  already  been  foreshadowed  in  enumerating  the 
drawbacks  connected  with  the  present  system,  which  becomes 
particularly  inconvenient  when  converters  are  employed.  I  reca- 
pitulate briefly. 

The  chief  advantages  that  may  be  wholly  or  partially  gained  by 
the  use  of  a  reverberatory  forehearth  in  works  where  the  blast- 
furnace matte  goes  to  Bessemer  converters  are: 


1.  The  saving  of  the  remelting-cupola  operation. 

2.  The  reduction  of  fnel  in  the  blast-furnace  to  its  lowest  limits, 
as  the  ore  requires  no  more  heat  than  is  sufficient  to  melt  it  so  that 
it  will  run  out  of  the  furnace. 

3.  The  complete  escape  from  all  the  delays  and  costs  connected 
with  the  chilling-up  and  burning-out  of  forehearths. 

■4.  An  increasing,  rather  than  a  diminishing,  temperature  as  the 
slag  flows  tlirough  the  settling  device.  This  is,  naturally,  a  most 
favorable  circumstance  for  the  separation  of  the  matte.  The  set- 
tling is  also  favored  by  the  constant  presence  of  a  large  body  of 
very  hot  matte  in  the  forehearth. 

5.  The  guarantee  of  any  desired  amount  of  matte  for  the  con- 
verters at  a  moment's  notice. 

6.  Permits  irregular  running,  or  even  a  complete  stopping  of 
the  converters-,  without  embarrassing  the  blast-furnace  work ;  for 
it  is  as  easy  to  tap  the  excess  of  matte  into  a  sand  bed  as  into  the 
converters,  and  when  the  latter  needs  more  matte  than  is  furnished 
by  the  ore,  the  pigs  can  be  slowly  charged  back  direct  into  the 
reverberatory,  and  melted  down  without  any  extra  fuel,  their  great 
fusibility  and  conductivity  making  this  possible. 

7.  By  keeping  a  stock  of  extra  rich,  and  extra  poor  matte  on 
hand,  and  charging  the  one  or  the  otlier  direct  into  the  partly 
drained  reverberatory,  the  grade  of  the  converter  charges  can  be 
rapidly  varied.  % 


These  important  points  are  discussed  in  the  chapter  on  "  Pyritie 
Smelting,"  but  I  desire  to  supplement  the  same  by  a  few  words 
regarding  the  subject  when  considered  from  the  standpoint  of 
ordinary  blast-furnace  smelting. 

With  our  present  knowledge,  it  ^eems  to  me  that  blast-furnaces, 
whether  water-cooled  or  of  brick,  fall  naturally  into  two  classes: 

1.  Blast-furnaces  used  simply  for  melting-down  ores  or  other 

2.  Blast-furnaces  used  for  partial  oxidation  as  well  as  melting. 
1.  Blast-furnaces  used  simply  for  melting-doivn  ores  or  other 


But  a  small  proportion  of  the  copper  blast-furnaces  of  the  world 
fall  strictly  within  this  category.  Typical  examples  of  such  fur- 
naces may  be  found  in  ordinary  foundry  cupolas  for  the  remelting 
of  pig  iron  for  castings,  or  in  the  cupolas  for  remelting  matte  for 


our  copper  Bessemer  couverters.  These  examples  are  particularly 
striking  because  the  materials  treated  are  free  from  gangue  and 
from  volatile  coustituents,  and  consequently  yield  (practically 
speaking)  no  slag,  nor  is  their  weight  diminished,  or  their  value 
increased,  by  the  operation.  The  furnace  process  produces  no 
chemical  action  in  the  charge.  It  merely  changes  the  substances 
into  a  more  convenient  form  for  future  treatment.  As  there  is 
no  chemical  change  in  the  ore,  it  follows  that  all  the  heat  neces- 
sary for  its  fusion  must  be  derived  from  coke,  or  other  extraneous 
fuel.  It  is,  therefore,  a  peculiarly  wasteful  and  unsatisfactory 
operation,  and  after  costing  a  considerable  sum  for  labor,  fuel, 
plant,  time,  and  metal-losses,  has  not  improved  the  actual  condi- 
tion of  the  substances  treated  by  a  single  iota. 

But  there  is  a  larger  class  of  operations  where  better  results  are 
obtained  by  this  same  neutral,  or  reducing,  system  of  smelting. 
This  is  where  a  certain  proportion  of  the  constituents  of  the  ore 
are  volatile,  or,  still  more  where  they  consist  partly  of  oxides  and 
silica  (gangue).  In  the  case  of  the  volatile  constituents,  we  re- 
move these  by  merely  melting  the  ore  as  already  explained,  and 
thus  effect  a  certain  slight  concentration,  the  value  of  the  product 
being  in  direct  proportion  to  the  amount  of  its  volatile  constitu- 
ents. To  take  an  extreme  case:  Suppose  our  ore  to  consist  of 
pure  iron  pyrites  without  gangue,  and  carrying  10  ounces  silver  to 
the  ton.  Iron  pyrites  contains  53  per  cent,  sulphur,  one-half  of 
which  is  so  loosely  bound  that  it  volatilizes  as  metallic  sulphur  at 
a  moderate  heat.  Hence,  100  pounds  of  this  ore  would  yield,  on 
fusion,  only  734^  pounds  of  product,  that  would  contain  silver  at 
the  rate  of  13. G  ounces  per  ton.  This  illustrates  a  concentration 
effected  by  volatilizing  certain  of  the  valueless  portions  of  the  ore. 

A  concentration  brought  about  by  causing  the  already  oxidized 
bases  of  the  ore  to  unite  with  the  silica  present,  is  a  much  more 
common  and  more  effective  operation. 

A  typical  example  may  be  found  in  the  Mansfeld  practice,  where 
the  ore,  as  it  comes  to  the  furnace,  contains  nothing  volatile,  and 
the  smelting  operation  is  conducted  in  so  powerfully  reducing  an 
atmosphere  that  none  of  the  coustituents  of  the  ore  can  become 
oxidized  in  the  furnace.  The  chemical  action  in  the  blast-furnace 
is  here  couBned  to  the  uniting  of  the  silica  with  such  bases  as  are 
already  oxidized.  But  as  the  ore  consists  mainly  of  silica,  mag- 
nesia, lime,  and  alumina,  with  a  little  oxidized  iron,  a  high  degree 
of  concentration  is  obtained   by  this  single  reducing  fusion,  a  45r 


per  cent,  matte  being  produced  from  a  3  per  cent,  ore,  while  some 
15  tons  of  slag  go  over  the  dump  for  each  ton  of  product. 

This  is  a  unique  case,  for,  as  a  rule,  our  ores  contain  so  large  a 
proportion  of  sulphur  as  sulphide  of  iron,  or  other  sulphides,  which 
will  combine  in  the  blast-furnace  with  already  oxidized  iron,  and 
steal  it  from  the  slag,  where  it  is  needed,  to  carry  it  into  the 
matte,  and  thus  augment  the  quantity,  and  decrease  the  quality, 
of  that  product,  that  it  is  customary  to  roast  the  ore,  by  which 
process  much  of  the  sulphur  burns  away  as  sulphurous  acid  gas,  and 
the  iron  is  oxidized  so  that  it  can  unite  with  the  silica  to  form  slag. 
We  thus  change  a  highly  pyritous  ore  to  a  condition  in  which  it 
somewhat  resembles  the  Mansfeld  ore.  That  is,  we  alter  it  so 
that  it  shall  consist  of  a  minute  proportion  of  metal  (sulphides), 
and  an  overwhelming  amount  of  gangue  rock  (oxides),  for  iron, 
when  oxidized,  may  be  regarded  as  gangue.  If  this  alteration  is 
sufficiently  thorough,  our  simple  reducing  smelting  will  bring 
about  the  desired  result:  i.e.,  a  small  proportion  of  rich  matte 
and  a  large  proportion  of  fusible  and  poor  slag. 

But  a  calcination  so  thorough  as  to  accomplish  this  result  is 
expensive  and  not  always  practicable,  for  many  ores  contain  too 
little  sulphur  to  warrant  roasting,  while  they  have  too  much  sul- 
phur to  yield  a  rich  matte  if  simply  melted  down  in  a  reducing 
cupola.  Leaving  out  extra  rich  ores,  it  may  be  said  that  three- 
fourths  of  all  the  copper  mines  in  the  world  are  able  by  ordinary 
roasting  and  reducing  smelting  to  produce  a  30  per  cent,  or  35  per 
cent,  matte  from  their  average  ores.  But  in  the  light  of  our  pres- 
ent practice,  this  is  an  exceedingly  inconvenient  product.  It  lacks 
some  15  per  cent,  of  being  rich  enough  to  send  to  the  converters, 
while  it  is  too  rich  to  make  it  advantageous  to  crush  and  calcine  it 
for  a  concentration  smelting.  It  is  the  mission  of  the  second 
division  of  blast-furnaces  to  add  this  lacking  15  per  cent,  of  copper, 
without  any  additional  operation. 

Blast-furnaces  that  are  used  simply  for  melting,  without  any 
desire  to  oxidize  the  charge  and  thus  enrich  the  matte,  are  charac- 
terized chiefly  by  the  following  features: 

(rt)  Contraction  toward  the  tuyeres  (boshes). 

(h)  High  ore  column. 

(c)  Strong  blast  pressure  (rapid  smelting). 

{d)  Small,  or  moderate-sized  tuyeres, 

(e)  Hot  blast.  (This  is  not  a  common  adjunct,  but  would  prob- 
ably always  be  economical  and  effective  for  this  peculiar  class  of 


I  have  spoken  hitherto  as  though  this  simple  reducing  smelting 
were  only  in  place  under  two  conditions: 

{a)  For  the  mere  object  of  changing  the  form  of  materials,  as 
in  melting  pig  iron  for  casting,  or  remelting  matte,  or  rich  ores, 
for  the  converters. 

(b)  For  smelting  ores  that  consist  mainly  of  silica  and  bases  in 
an  oxidized  condition  (either  naturally  or  by  roasting). 

I  am  strongly  of  the  opinion  that  a  third  condition  may  soon  be 
added  to  these,  the  success  and  economy  of  modern  converter  work 
having  greatly  changed  the  relation  of  the  various  metallurgical 
processes  to  each  other. 

At  present,  in  America,  we  do  not  like  to  bessemerize  matte 
that  runs  very  much  below  50  per  cent,  copper,  45  per  cent,  being 
the  extreme  limit  for  regular  work.  It  would  be  considered  ridic- 
ulous to  attempt  to  bessemerize  a  matte  containing  only  20  per 
cent,  or  even  15  per  cent,  copper.  There  are  three  main  ditiicui- 
ties  in  the  way  of  effecting  this  exceedingly  desirable  object: 

1.  Converter  linings  become  too  rapidly  destroyed  by  mattes 
below  45  per  cent,  copper,  and  no  basic,  or  artificially  cooled,  lining 
has  yet  been  a  success,  nor  have  we  been  able  to  induce  the  ferrous 
oxide  produced  from  the  matte  to  content  itself  with  artificially 
supplied  silica  instead  of  robbing  it  from  the  lining. 

2.  Slag  is  made  too  rapidly  when  the  matte  contains  much  iron, 
and  no  method  for  its  continuous  removal  from  the  converter  has 
yet  been  successful. 

3.  The  amocint  of  copjDer,  or  of  rich  matte,  derived  from  a  very 
low  grade  matte  is  too  small  to  manipulate  without  some  continuous 
method  of  introducing  fresh  matte. 

If  these  difficulties  were  obviated,  and  none  of  them  appear  in- 
superable, it  seems  to  me  that  where  coal  is  cheap  and  coke  dear, 
as  in  many  places  in  the  "West;  or  where  water-power  is  available, 
as  at  Great  Falls,  Montana,  our  simplest  and  most  economical  way 
of  handling  such  ores  as  those  of  Butte  (or  of  most  other  American 
copper,  and  copper-silver-gold  districts),  will  be  to  smelt  them  raw 
in  large  blast-furnaces  with  coke  and  a  hot  blast,  creating  a  power- 
ful reducing  action,  and  running  the  low-grade  matte  continuously 
into  Bessemer  converters,  where  it  will  be  blown  up  to  a  point 
when  the  resulting  slag  becomes  rich  enough  to  require  resmelting, 
(vvhich,  with  reverberatory  settlers  may  be  60  per  cent,  or  more). 
This  matte,  tapped,  or  run  direct  into  the  finishing  converters, 
wiU  yield  a  very  small  amount  of  slag  for  re-treatment,  the  operation 


being  so  regulated  that  there  will  be  just  enough  converter  slag  to 
flux  the  highly  siliceous  raw  ore  in  the  blast-furnaces.  I  would 
propose  to  greatly  contract  the  present  processes  of  mechanical 
concentration  at  Butte,  and  a  very  small  proportion  of  the  copper 
thus  rescued  from  loss  would  pay  for  the  extra  coke  required  to 
suielt  the  raw  ore.  The  ore  slags  might  easily  run  from  45  per 
cent,  to  50  per  cent,  silica,  and  would  be  specifically  very  light, 
and  contain  under  0.3  per  cent,  of  copper.  This  would  greatly 
simplify  and  cheapen  the  entire  metallurgical  plant  and  treatment, 
and,  in  the  instance  specified,  would  largely  substitute  the  power 
of  the  Missouri  River  for  hand  labor  and  fuel.  It  would  abolish 
the  crushing  and  roasting  of  the  ore  and  curtail  the  process  of 
mechanical  concentration  by  some  60  per  cent,  or  more. 

The  Butte  metallurgists  have  faced  and  solved  problems  consid- 
erably more  difficult  than  this  one  appears  to  be.  The  bessemeriz- 
ing  of  matte  containing  20  per  cent.,  and  less,  of  copper  is  an 
accomplished  fact  in  France  and  Russia,  though  I  have  not  myself 
seen  it,  nor  do  letters  to  me  from  metallurgists  engaged  in  the 
work  give  me  any  satisfactory  practical  reasons  of  how  they  induce 
linings  to  stand  under  such  circumstances.  As  regards  the  chang- 
ing of  the  converter  process  from  an  intermitteutj  to  a  continuous, 
operation,  I  cannot  see  that  any  insuperable  obstacle  exists. 

2.  Blast-furnaces  for  partial  oxidation  as  well  as  for  melting. 

This  section  comprises  by  far  the  greater  portion  of  the  copper 
blast-furnaces  of  thisj  and  ocher  countries.  The  operation  varies 
from  a  slight,  and  often  unsuspected,  oxidation  of  a  little  of  the 
sulphur  and  iron  of  the  charge  when  smelting  ordinary  raw  or 
roasted  ore,  to  the  most  pronounced  form  of  pyritic  smelting. 

As  this  latter  process  is  considered  fully  in  a  separate  chapter,  I 
must  confine  myself,  in  this  section,  to  furnaces  where  no  especial 
attempt  is  made  to  utilize  the  ore  itself  as  fuel,  or,  in  other  words, 
to  practise  pyritic  smelting. 

The  extent  to  which  oxidation  shall  be  pushed  in  the  blast-fur- 
nace is  a  point  that  has  a  most  important  bearing  on  the  economy 
of  the  entire  process,  and  one  that  demands  for  its  correct  decision 
the  greatest  experience  and  judgment  on  the  part  of  the  metallur- 
gist. Each  case  has  to  be  judged  upon  its  own  merits;  but  under 
the  great  majority  of  conditions  and  with  the  present  general 
arrangement  and  construction  of  plants,  it  will  be  found  decidedly 
advantageous  to  use  the  blast-furnace  as  a  partial  oxidizer,  and  to 
produce  a  richer  matte  than  would  naturally  result  if  the  charge 


were  simply  melted  down  in  a  reducing  smelting,  as  occurs  in  the 
crucible  assay  for  determining  the  amount  of  matte  that  will  be 
produced  by  a  given  mixture. 

Xo  one  need  shrink  from  this  practice  as  a  dangerous  or  untried 
experiment.  Probably  the  very  metallurgist  who  would  refuse  to 
listen  to  a  suggestion  to  use  his  blast-furnace  as  a  partial  roaster  or 
calciner,  is  actually  running  it  more  or  less  on  these  lines  without 
ever  having  realized  the  fact.  If  he  doubt  the  truth  of  this  state- 
ment, let  him  merely  decrease  the  size  of  his  hearth  and  of  the 
shaft  slightly  above  the  tuyeres,  use  smaller  tuyeres,  and  thinner 
layers  ct  charge,  and  a  stronger  blast.  Then,  when  he  observes 
his  matte  increase  in  quantity  and  decrease  iu  quality,  and  his 
slag  become  siliceous  from  the  robbery  of  its  iron  by  the  unoxidized 
sulphur,  he  will  realize  that  he  has  been  partially  calcining  his  ore 
in  his  blast-furnace,  and  has  been  practising  what  I  term  "Com- 
promise Pyritic  Smelting." 

It  is  frequently  a  matter  of  the  greatest  value  to  employ  this 
partial  oxidation  of  the  charge  in  the  blast-furnace,  and  it  is  always 
useful  to  feel  that  one  at  least  knows  how  to  accomplish  it  if  occa- 
sion should  require  it. 

The  difference  between  this  method  and  the  rapid  process  of 
merely  melting  the  ore,  which  was  considered  in  the  previous  sec- 
tion, lies  entirely  in  so  running  the  furnace  that  a  partial  oxidizing 
atmosphere  is  substituted  for  the  powerful  reducing  atmosphere 
that  characterizes  the  other  operation.  This  is  effected  mainly  by 
ditiusing  the  heat  over  a  greater  area  and  lessening  the  sudden 
violence  of  the  combustion  at  the  tuyeres.  When  we  wished  to 
simply  melt  the  ore  in  the  most  rapid  manner  possible,  we  con- 
stricted the  shaft  at  the  tuyeres  and  blew  a  strong  blast  into  this 
concentrated  mass  of  coke  and  ore,  producing  a  very  high  local 
temperature  and  a  dense  atmosphere  of  carbonic  oxide  gas.  The 
ore  melted  almost  instantaneouslv  and  dropped  into  the  neutral 
hearth  below.  The  sinking  of  the  charge  was  rapid,  the  heat  was 
concentrated  in  the  tuyere  zone,  and  the  ore  had  scarcely  reached 
a  red  heat  before  it  was  fused  and  removed  from  all  chemical 

To  obtain  a  certain  amount  of  oxidizing  effect,  we  need  pretty 
much  the  opposite  set  of  conditions,  and  the  mere  enumeration  of 
one  or  two  of  them  suggests,  or  rather  compels  the  remainder. 
We  need  a  light  blast;  but  a  light  blast  cannot  penetrate  a  thick 
column  of  charge,  nor  will  it  give  any  reasonable  capacity  for  the 


furnace.  We  are  forced,  tlierefore,  to  use  a  furnace  which  is  nar- 
row in  one  of  its  dimensions,  so  that  the  blast  can  penetrate  the 
ore  column,  and  we  must  lengthen  it  in  the  other  direction  in 
order  to  obtain  sufficient  capacity.  This  brings  us  to  the  long, 
narrow  rectangle  as  the  only  suitable  form  for  our  purpose,  and 
furnaces  are  now  constructed  with  a  shaft  up  to  14  feet  in  length, 
the  ordinary  width  being  32  inches  to  38  inches.  We  desire  to 
avoid  the  concentration  of  heat  and  the  reducing  effect  inseparable 
from  a  contraction  of  the  shaft  at  the  tuyeres,  and  find  that  we 
obtain  the  best  oxidizing  results  from  perfectly  perpendicular  walls. 
As  the  blast  pressure  must  be  light,  we  make  up  our  deficiency  in 
oxygen  by  increased  volume  of  wind,  and  consequently  are  obliged 
to  enlarge  the  diameter  of  our  tuyeres  to  -i  inches  and  even  6 
inches.  A  high  ore  column  strongly  favors  reduction.  Hence, 
we  employ  an  ore  column  only  high  enough  to  utilize  the  heat  as 
far  as  practicable,  and  to  give  the  ore  time  for  partial  oxidation 
during  its  descent.  Four  to  six  feet  from  tuyeres  to  charge  door 
is  the  average  height.  We  expect  our  charge  to  be  moderately  hot 
on  top,  as  the  furnace  is  acting  as  a  roaster  to  the  very  tunnel- 
heald.  We  prefer  a  cold  blast,  as  heated  wind  leads  to  the  concen- 
tration of  temperature  and  rapid  smelting  that  we  are  trying  to 

To  obtain  the  large  volume  and  low  pressure  of  blast  that  we  re- 
quire, a  fan  blower  may  quite  possibly  be  the  most  effective  and 
economical  machine  that  we  can  employ.  Its  main  disadvantages 
are  its  high  speed,  small  pulleys,  and  large  belts. 


The  char  (J  iiuj  of  the  llast- furnace  by  shovel  is  being  gradually 
replaced  by  more  or  less  perfect  mechanical  devices.  Where  hand 
labor  is  still  employed,  the  ore  and  coke  should  flow  from  bins 
direct  into  two-wheeled  charging  barrows,  that  can  be  dumped 
upon  the  cast-iron  floor  at  the  charging  door  of  the  furnace. 
Scoop  shovels  should  be  used  in  charging  both  ore  and  coke,  and 
no  man  who  finds  them  too  heavy  can  make  a  rapid  feeder.  The 
railroad  dump-cars  will,  of  course,  run  directly  over  the  charging 
bins,  and  drop  their  contents  into  the  latter.  Where  the  lay  of 
the  ground  is  unsuitable  for  a  terraced  construction,  an  inclined 
plane  with  winding-engine  to  haul  the  railroad  cars  up  over  the 
bins  is  much  more  economical  than  any  form  of  elevator,  and  much 
better  suited  to  handling  large  quantities  of  material  without  con- 


fiisioD.     All  ore,  coke,  and  slag  should  be  delivered  in  this  mannei 
and  wheelbarrows  should  be  regarded  with  susj)icion. 

It  is  cheaper  and  more  convenient  to  dump  fuel  and  ore  into  the 
furnace  direct  from  the  charging  barrows.  To  do  this  to  advan- 
tage,  it  is  necessary  to  construct  the  blast-furnace  with  aside  flue 
through  which  the  gases  are  drawn  off  below  the  level  of  the 
charging  door.  This  is  arranged  in  the  same  manner  as  with  lead 
furnaces,  by  a  thimble  introduced  into  the  upper  portion  of  the 
shaft,  the  gases  being  drawn  off  from  the  annular  space  between 
the  thimble  and  the  furnace  walls.  By  this  device,  the  bulky 
housings  and  overhead  flue  are  abolished,  and  the  furnace  opening 
consists  merely  of  a  rectangular  slot  in  the  unencumbered  charging 

The  Pueblo  Smelting  Company  has  adapted  an  excellent  device 
whereby  the  filling  of  the  furnaces  is  accomplished  by  means  of  a 
long,  narrow  charging-car  corresponding  to  the  rectangular  open- 
ing of  the  furnace-top,  and  running  on  a  track  that  straddles,  and 
is  at  right  angles  to  the  long  axis  of  all  the  blast-furnace  tunnel- 
heads.  By  a  simple  stop-mechanism,  the  attendant  controls  the 
car  so  that  it  shall  deliver  its  load  of  coke  and  ore  into  any  furnace 
requiring  it.  A  single  man  on  the  charging  floor  can  thus  attend 
to  the  charging  of  six  large  furnaces. 

So  far  from  finding  it  derange  the  running  of  the  furnace,  I 
have  obtained  better  and  more  uniform  results  the  nearer  I  have 
approached  to  strictly  automatic  charging.  If  one  corner  of  the 
furnace  threatens  to  chill,  it  is  easy  to  arrange  the  mechanical 
device  so  the  ore  shall  be  diverted  from  the  chilled  portion  for  a 
charge  or  two,  and  the  substitution  therefor  of  a  few  hundred- 
weight of  basic  slag,  and  the  plugging  of  the  one  or  two  tuyeres 
that  are  involved  in  the  chill,  will  soon  set  matters  right  again. 

Tlie  Itandhng  of  the  prochicts  of  the  Uad-furnnce  has  also  been 
considerably  cheapened  of  late  years. 

The  matte  is  either  tapped  off  at  intervals  into  slag-pots,  only 
about  1,000  pounds  being  drawn  off  at  each  tapping,  in  order  that 
the  matte  in  the  forehearth  may  not  be  unduly  lowered,  or  it  is 
tapped  in  large  charges  direct  into  the  converters,  or  converter- 
ladles,  or  it  may  be  tapped  in  considerable  amounts  into  sand  bed& 
or  iron  (or  soft  steel)  molds.  The  latter  method  is  generally  used 
where  matte  is  to  be  shipped  or  sold,  as  it  gives  a  cleaner  product 
and  lessens  the  chance  of  irregularities  in  the  sampling.  Jt  can 
also  easily  be  shotted  by  a  strong  jet  of  water,  though  this  makes. 

blast-fuk>;ace  smeltixg.  313 

mauy  polished,  bean-like  granules  that  seem  scarcely  worth  the 
trouble  of  crushing,  and  yet  resist  the  action  of  the  calcining 

Slag  may  be  handled 

1.  In  small  i)ots  by  man  power. 

2.  In  large  pots,  or  other  vessels,  by  mule,  or  steam  power. 

3.  In  mechanical  pan  conveyers, 
•i.   By  granulation  by  water. 

1.  In  small  slag-pots.  Although  these  useful  little  pots  are 
being  lapidly  superseded  by  more  economical  devices,  they  still  re- 
tain their  place  at  many  good  works,  and  are  worthy  of  careful 

Mr.  H.  A.  Keller,  in  an  excellent  paper  on  the  subject,*  gives 
some  interesting  cuts  of  slag-pots,  which  I  copy,  together  with 
his  description  and  comments. 

'*Iu  the  accompanying  illustrations.  Figs.  2  to  5  inclusive  repre- 
sent the  cart  now  in  use  at  the  Parrot  works,  and  Figs.  6  to  9  a 
cart  similar  to  the  one  introduced  by  the  writer  at  the  Philadelpliia 
works  at  Pueblo,  Colorado. f  Parts  of  these  pots  have  been  in  use 
for  a  number  of  years,  while  other  parts  are  of  more  recent  date. 

The  cast-iron  track  shown  in  the  drawings  is  laid  into  that  part 
of  the  slag-dump  which  by  constant  usage  is  apt  to  become  spe- 
cially rough  and  uneven.  A  rough  dump,  besides  adding  to  the 
work  of  the  slag-wheeler,  greatly  increases  the  necessary  repairs. 
Further  away  from  the  furnaces  the  dump  is  leveled  by  "slag 
squares"  or  slabs  of  slag  formed  by  pouring,  which  are  constantly 
kept  up  to  its  edge.  These  are  best  made  2  feet  by  4  feet  and 
from  8  to  12  inches  deep.  After  a  mold  of  these  dimensions  has 
been  formed  by  means  of  rails  or  cast-iron  plates  and  cold  slag,  it 
is  partially  filled  with  large  pieces  of  cold  slag,  which  are  then 
cemented  together  with  liquid  slag  poured  simultaneously  from 
several  pots. 

The  slag-cart  consists  of  three  parts,  the  bowl  or  pot  proper,  the 
wheels,  and  the  handle  or  foot. 

There  are  two  styles  of  bowls  now  in  general  use.  These  are 
represented  in  the  accompanying  Figures,  as  No.  1  and  No.  2. 
On  account  of  its  straight  sloping  sides,  the  pointed  bowl.  No.  2, 
allows  the  matte  to  settle  more  readily.     It  is  therefore  preferred 

*  Transactions  American  Institute  Mining  Engineers,  Vol.  XXII.,  p.  574. 
+  Hofman's  Metallurgy  of  Lead,  p.  203. 


when  but  little  matte  is  suspeuded  in  the  liquid  slag,  which  matte 
is  to  be  saved  iu  shell  and  bottom.  To  avoid  unnecessary  dis- 
turbance, this  bowl  is  provided  with  a  1^-inch  hole,  through  which 
the  liquid  slag  is  tapped  by  a  ^-inch  bar,  usually  of  hexagonal 
steel.  Such  a  tap-hole  was  first  used  in  this  country  by  Mr.  W. 
B.  Devereaux,  at  Aspen,  Colorado.*  Its  location  varies,  of  course, 
with  circumstances.  After  much  experimenting,  the  writer,  for 
instance,  when  employed  at  the  Philadelphia  works,  determined  to 
locate  it  as  shown  in  Fig  8. 

The  bowl  No.  1,  witli  rounded  sides,  has  the  advantage  of  greater 
capacity  than  one  of  similar  dimensions  with  straight  sides.  Such 
a  bowl  is  therefore  preferable  when  it  is  intended  to  dump  out  the 
entire  cone.  A  more  shallow  bowl  (shown  in  Fig.  1),  introduced 
by  Mr.  A.  Filers  and  universally  used  in  early  Leadville  smelting 
practice,  is  gradually  disappearing  with  the  increased  size  of  slag- 
dumps,  since  it  does  not  permit  as  great  capacity  as  those  shown 
in  Figs.  2  to  9. 

Bowl  No.  1  requires  the  axles  to  be  fastened  to  it  with  set- 
screws,  while  the  straight  sides  of  No.  2  leave  room  for  securing 
the  axles  with  wedges.  In  the  latter  case,  each  axle  is  provided 
with  a  square  stub  :j-inch  larger  than  the  diameter  of  the  axle. 
The  axles  of  the  third  form  of  bowl  (Fig.  1),  are  carried  by  one 
continuous  piece  of  square  iron,  bent  to  conform  with  its  shape. 
This  is  fastened  to  the  solidly  cast  side-lugs  by  stirrup-clamps  and 
further  held  in  place  by  passing  between  two  guiding-lugs  at 
bottom  of  bowl. 

The  lug  shown  at  the  rim  of  each  bowl  is  a  great  protection  to 
the  spot  where  dumpiug  causes  most  wear.  It  was  first  suggested 
by  Mr.  0.  T.  Limberg  of  Leadville,  and  is  now  universally  used. 

The  splash-guards  prevent  the  liquid  slag  from  spilling  upon 
the  hubs.  They  were,  I  believe,  first  introduced  at  the  Grant 
works,  Denver,  Colorado. 

The  false  bottom  for  pot  No.  1,  shown  in  dotted  lines  in  Figs. 
2  and  4,  has  been  in  use  for  several  years  at  the  Parrot  works.  It 
consists  of  a  cast-iron  disc  held  in  place  by  a  countersunk  |-incli 
bolt.  If  the  original  bottom  is  very  badly  damaged,  a  washer  may 
have  to  be  used  besides.  Though  such  an  arrangement  does  not 
give  satisfaction  with  a  heavy  flush,  it  behaves  admirably  with  slag 
or  matte  running  slowly.     With  matte,  it  has  the  additional  ad- 

*  See  Keri's  MetallMlHenkunde.  1881,  p.  100. 


vantage  of  producing  a  flattened  cone  which  is  more  easily  broken 
than  a  tapering  one.  New  bowls  are  used  for  slag  at  the  Parrot, 
and  last  about  eighteen  months.  After  that,  being  provided  with 
these  false  bottoms,  they  last  almost  as  long  on  matte.  Another 
contrivance  for  usiug  a  bowl  after  its  point  is  worn  out  has  been 
described  by  Mr.  R.  H.  Terhune.* 

By  the  device  shown  in  No.  1,  the  bowl  is  made  reversible,  the 
cart  being  at  the  same  time  steadied  by  fastening  the  handle  with 
two  straps  instead  of  one. 

The  style  of  wheel  vepresentedf  consists  of  a  cast-iron  hub,  with 
wrought-iron  spokes  and  tire.  It  is  mounted  upon  a  steel  roller- 
bearing.  The  hull  is  tapped  to  receive  the  end  of  the  spoke,  which 
for  that  purpose  is  threaded.  For  further  tightening,  each  spoke 
is  provided  with  a  Jam-nut.  After  the  spokes  have  beeu  carefully 
adjusted,  the  wrought-iron  tire  is  shrunk  upon  their  outer  ends 
and  is  subsequently  fastened  to  them  by  means  of  countersunk 
rivets.  This  tire  being  the  part  of  the  wheel  subjected  to  most 
wear  must  be  made  sufficiently  heavy  and  strong  without  being 
clumsy.  The  advantages  of  a  wrought  over  a  cast  tire  are  evident, 
particularly  when,  as  in  this  case,  the  former  is  well  fastened  and 
easily  repaired.  The  anti-friction  rollers  require  no  oiling,  or  at 
least  but  little.  A  double  set  of  these  rollers  is  put  in  loosely 
around  each  axle.  Thus  arranged,  they  are  prevented  from  wear- 
ing so  as  to  cause  appreciable  friction,  i.  e.,  by  running  upon  one 
another.  By  using  f-iuch  rollers,  a  l^-inch  axle  takes  two  sets  of 
nine;  and  for  each  ^-inch  increase  of  axle-diameter  an  additional 
roller  is  required. 

The  foot  or  handle  is  practically  the  same  for  all  forms  of  slag- 
cart.  The  essential  point  is  that  it  shall  be  of  sufficient  length, 
and  its  crosspiece  wide  enough  for  convenient  pushing.  It  is  fully 
illustrated  in  the  drawings  and  needs  no  further  comment. 

All  rivets  and  bolts  in  either  No.  1  or  No  2  are  f-inch  in  diam- 
eter. The  average  weight  of  No.  1  is  563  pounds.  Many  matte- 
cones  taken  from  such  pots  with  false  bottoms  gave: 

631  pounds  for  55  per  cent,  copper. 

603^  pounds  for  52  per  cent,  copper. 

597  pounds  for  47  per  cent,  copper. 

574^  pounds  for  44  per  cent,  copper. 

*  Trans.,  XV.  92.  +  Patented  by  Cole,  Gaylord  and  Keller^ 


A  large  numlier  of  slag-cones  taken  from  pots  with  original  bot- 
toms gave  an  average  of  -iOG  pounds.     Their  composition  was: 

SiO,  38 

FeO    46.5 

A1,0. 10.5 

CaO 3 

Total 98 

The  copper  contents  are  from  0.3  to  0.4  per  cent,,  present  either 
as  CiijO  or  as  suspended  matte. 

It  may  be  of  interest  to  mention  here  that  a  large  number  of 
copper-matte  analyses  from  reverberatory  and  blast-furnaces,  com- 
prising the  diflEerent  grades  and  extending  over  several  years,  gave 
a  constant  tenor  of  from  21  to  23  per  cent,  sulphur.  Many  of 
these  analyses  showed  the  presence  of  magnetic  iron,  sometimes  in 
considerable  quantity.  Copper-mattes  would  accordingly  corre- 
spond to  either  of  these  formul^t:* 

{CuS)r  +  (FeoS)y  or 

(Cn,S),  +  (Fe8),  -f  (FeA).- 

No.  1  slag-pot  casts,  as  a  rule,  six  full  slag-bricks.  A  great 
many  bricks,  being  weighed,  gave  an  average  weight  of  54  pounds 
each.  The  slag  given  above  weighs  therefore  216  pounds  per  cubic 
foot.  Taking  the  same  quantity  of  water  at  62^  pounds,  gives  a 
specific  gravity  of  3.47.  These  seem  fully  as  accurate  as  similar 
determinations  made  in  many  laboratories  of  Western  smelting 
works,  most  of  which  are  necessarily  crude  in  this  line  of  research." 

The  foregoing  remarks  indicate  the.  chief  points  that  require 
attention  in  the  construction  of  slag-pots.  If  they  are  properly 
made,  there  is  one  certain  way  to  make  them  last.  This  is  to 
liave  plenty  of  them,  so  that  they  may  have  a  chance  to  cool  before 
being  used  again.  The  damage  effected  by  constant  overheating 
is  far  beyond  the  mere  burning  of  the  iron  or  distortion  of  the 
pot.  The  main  injury  arises  from  the  slag  or  matte  **  welding" 
to  the  overheated  pot  and  requiring  much  hard  sledging  on  the 
exterior  before  the  slag  cone  will  drop  out  of  it.  It  is  this  treat- 
ment that  destroys  pots. 

Another  point  in  regard  to  the  economical  handling  of  slag  in 
these  pots  is  the  importance  of  extreme  neatness  about  the  dump, 

*The  writer  is  indebted  to  Dr.  Edward  Keller  for  these  formulae. 


furuace,  and  runway.  An  old  pot-hauler,  who  understands  his 
business,  will  keep  his  floor  and  runway  so  free  from  splashes  of 
slag  that  lie  can  push  his  pot  with  the  very  slightest  expenditure 
of  power.  He  keeps  three  pots  always  at  the  furuace;  a  pot  that 
Has  just  been  filled  and  is  waiting  until  a  thin  skin  has  formed  on 
the  surface  to  prevent  splashing;  a  pot  that  is  being  filled;  and  a 
cold,  empty  pot  to  use  next.  When  the  pot  at  the  furnace  is 
about  five-sixths  full,  the  furnace-man  holds  a  ladle  under  the  slag 
stream,  while  the  pot-hauler  removes  the  full  pot  and  shoves  in  the 
empty  one.  The  full  pot  is  allowed  to  stand  at  one  side,  in  order 
to  chill  on  the  surface,  while  the  one  that  is  already  skimmed  over 
on  top  is  pushed  oat  on  to  the  dump.  This  skimming  over  of  the 
surface  prevents  all  splashing  of  liquid  slag,  and  a  pot-hauler  who 
adopts  all  these  little  precautions  has  not  only  an  easy  track  to 
push  on,  but  has  almost  no  sweeping-up  to  do. 

It  is  equally  important  to  keep  the  brink  of  the  dump  in  good 
repair,  and  with  a  sharp  edge  and  steep  slope.  When  two  or  more 
furnaces  are  running,  a  dump-man  is  required  on  both  day  and 
night  shifts,  and  can  save  his  wages  several  times  over. 

2.  In  large  pots  or  vessels.  Dumps  have  grown  so  large,  and 
furnaces  smelt  so  much  more  ore  than  formerly,  that  it  has  been 
found  convenient  to  sling  two  or  more  large  pots  on  a  frame  run- 
ning on  a  track,  and  use  a  mule  to  drag  them  to  the  edge  of  the 
(lump.  The  pots  may  be  so  hung  as  to  be  easily  tipped,  or  their 
liquid  contents  may  be  tapped  through  a  hole  near  the  bottom. 
Large  frames  are  also  used,  which  stand  upon  an  iron  car,  and 
taper  a  little  toward  the  top.  When  the  enclosed  block  of  slag 
has  chilled  sufficiently,  the  frame  is  hoisted  off,  and  the  carriage 
is  inclined  by  suitable  mechanism  so  that  the  block  of  slag  slides 
off,  and  down  the  face  of  the  dump.  These  latter  carriages  may 
be  hauled  by  mules,  or  coupled  into  a  train  and  shifted  by  a  small 
locomotive.  Their  construction  and  manipulation  is  a  familiar 
part  of  the  metallurgy  of  iron. 

3.  Mechanical  pan-conveyers.  These  devices  also  have  been 
principally  developed  in  the  metallurgy  of  iron,  to  whose  text- 
books the  student  is  referred.  One  of  the  most  convenient  that  I 
know  is  an  English  patent  called  "Hawdon's  Slag  Carrier."  It  is 
plainly  shown  in  Fig.  43. 

4.  Granulation  by  water.  After  trying  various  shaped  jets  and 
other  more  or  less  elaborate  devices,  the  majority  of  the  smelters 
now  granulating  their  slag  by  water  liave  come  down  to  a  simnle 


stream  of  that  liquid,  runuing  through  a  narrow  trough  with  con- 

siderable velocity,  and  into  which  the  stream  of  slag  drops.     Very 
little  steam  or  noise  is  made,  and  the  practice  is  entirely  satisfactory 


providing  there  is  enough  water  to  thoroughly  granulate  and  re- 
move the  slag,  and  prevent  any  possible  formation  of  a  solid  cone 
with  liquid  contents,  which  might  cause  a  very  serious  explosion. 
I  know  of  no  accurate  figures  as  to  amount  and  pressure  of  water 
required  under  various  conditions;  but  I  have  found  that  a  stream 
of  water  delivered  by  a  two-inch  pipe  under  a  head  of  12  feet,  and 
flowing  through  a  launder  with  a  fall  of  one  inch  to  the  foot,  will 
thoroughly  granulate  and  remove  100  tons  of  heavy,  ferruginous 
slag  per  24  hours. 

One  of  the  main  difficulties  by  this  system  is  mechanical,  the 
destruction  of  the  launders  by  the  granulated  slag.  Cast-iron 
plates  are  generally  used  as  bottoms,  but  I  have  been  able  to  obviate 
the  expense  of  constant  renewals  by  forming  the  main  launder  of 
slag-brick.  The  bottom  will  last  a  long  time  and  can  be  quickly 

The  water  can,  of  course,  be  used  over  and  over,  a  waste  of  5 
per  cent,  being  experienced  at  each  time. 

Where  there  is  not  sufficient  fall  to  permit  of  the  direct  discharge 
of  the  slag  over  the  dump,  a  bucket  elevator  makes  an  ideal  ar- 
rangement for  elevating  the  granules  to  any  desired  height,  and 
thus  building  the  dump  up  in  the  air.  As  this  granulated  slag 
makes  excellent  material  for  roads  and  embankments,  the  sluice 
may  discharge  direct  into  the  railroad  cars,  the  water  leaking 
through  the  sides  of  the  car  or  flowing  over  the  top.  A  ditch  at 
the  lower  side  of  the  track  will  catc  h  all  the  water  and  lead  it  to  a 
suitable  reservoir. 

It  is  obvious  that,  when  granulating  the  slag  by  water,  a  careful 
watch  must  be  kept  on  the  level  of  the  matte  in  the  forehearth. 
An  excellent  control  can  be  kept  on  the  furnace-man  by  examining 
the  little  mound  at  the  foot  of  the  sluice,  as  any  matte  granules 
will  be  found  here  in  a  concentrated  form.  Besides  repeated  ex- 
aminations during  the  day,  this  mound  should  always  be  carefully 
panned  at  the  change  of  shift,  else  each  furnace-man  will  claim 
that  any  matte  found  therein  was  made  in  the  other  shift. 



Although  some  90  per  cent,  of  the  copper  blast-furuaces  in  the 
United  States  are  now  water-jacketed,  one  progressive  and  thor- 
oughly experienced  concern,  The  Orford  Copper  Company,  still 
uses  the  large  brick  Raschette  furnaces  that  they  introduced  some 
20  years  ago.  Such  results  as  they  obtain  in  these  furnaces,  which 
are  now  cooled  in  the  region  of  the  tuyere  openings  by  means  of 
water  circulating  in  pipes  embedded  in  the  brick  work,  cannot 
properly  be  ignored.  This  type  of  furnace  also  demands  so  much 
care  and  skill  in  its  management,  that  it  forms  a  peculiarly  instruct- 
ive study. 

The  distinctive  peculiarities  of  the  "Orford"  furnace,  as  this 
altered  and  improved  form  of  Raschette  furnace  is  usually  desig- 
nated, aside  from  its  unusual  size,  are  the  large  number  and  diam- 
eter of  its  tuyere  openings — 14  of  G  inches  diameter;  the  absence 
of  any  interior  crucible  or  space  for  the  collection  of  the  fused 
products;  the  substitution  therefor  of  an  exterior  forehearth  or 
basin,  and  the  construction  of  the  latter  in  such  a  manner  that 
two  continuous  streams — of  slag  and  metal  respectively— flow 
therefrom  into  ordinary  slag-pots,  without  any  blowing  through 
of  the  blast,  or  delay  for  tapping  and  other  related  manipulations.* 
The  latter  arrangement  may  be  applied  to  any  furnace  of  sufticient 
size,  it  being  absolutely  essential,  for  the  prevention  of  chilling, 
that  a  large  quantity  of  molten  material  should  constantly  traverse 
it.  If  the  product  is  a  matte  of  higb  grade,  00  per  cent.,  and 
over,  a  much  larger  quantity  is  necessary  to  prevent  chilling  than 
if  the  metal  is  of  poorer  quality.  The  rapid  chilling  of  the  foi-mer 
is  due  not  to  its  possessing  a  higher  fusion  point,  but  because  its 
capacity  as  a  conductor  of  heat  increases  with  its  percentage  of 

When  the  smelting  mixture  is  exceedingly  rich,  so  that  a  very 
large  amount  of  the  copper-bearing  product  results,  it  is  even  pos- 

*  See  section  on  "Foreheartlis  "  for  detailed  description  of  this  device. 



sible,  by  rapid  smelting,  to  maintain  a  constant  stream  of  metallic 
copper — '3  practice  that  may  be  regarded  as  a  curiosity  rather  than 
as  ordinarily  feasible. 

A  detailed  description  of  the  construction  and  subsequent  man- 
agement of  this  form  of  furnace  will  bring  forward  tlie  points 
already  referred  to,  and  illustrate  the  practice  that  up  to  the  pres- 
ent time  has  been  found  most  advantageous,  and  which  has  cheap- 
ened the  smelting  of  copper  ores  to  a  remarkable  extent. 

The  outside  measurement  of  the  furnace  being  8  feet  o  inches 
by  16  feet  8  inches,  an  excavation  should  be  made  at  the  intended 

Fig.  44. 


Orpord  Brick  Furnace. 

Fig.  45. 

site  some  three  feet  larger  in  every  direction  than  the  figures  just 
given,  and  of  sufficient  depth  to  reach  solid  ground  and  insure  a 
proper  foundation.  A  depth  of  4  or  5  feet  will  usually  suffice, 
the  pit  being  immediately  filled  with  concrete;  or,  where  possible, 
the  pit  should  be  filled  to  nearly  the  surface  with  molten  slag. 

The  walls  of  the  furnace  should  be  begun  a  foot  below  the 
ground  level,  and  should  consist  entirely  of  fire-brick  up  to  the 
tuyere  level,  where  the  panels  shown  in  the  cut  are  begun.  Up 
to  this  point,  the  walls  are  30  inches  thick,  of  solid  fire-brick, 
while  the  panels  are  only  18  inches  thick,  thus  being  more  accessi- 
ble for  repairs,  and  containing  the  tuyere  openings.  The  rear  wall 
is  divided  into  three  panels  equally  spaced,  and  supported  on  each 



side  by  the  full  thickness  of  the  wall,  forming  colnnins  at  each 
coruer,  aud  between  the  weaker  portions,  that  are  chiefly  relied 
upon  to  carry  the  weight  of  the  superincumbent  structure.  The 
panels  are  30  inches  wide  and  33  inches  high,  and  are  strongly 
arched  over  with  three  rows  of  fire-brick,  above  which  the  fall  thick- 
ness of  the  wall  (30  inches)  is  maintained  to  the  top  of  the  struc- 
ture. Each  panel  is  pierced  by  two  G-inch  square  tuyere-holes, 
equally  spaced,  excepting  the  central  front  panel,  which  contains 
only  a  small  orifice  for  the  slag-run,  at  a  point  some  10  inches 
below  the  tuyere  level.  The  panel  referred  to  forms  the  breast  of 
the  furnace,  and  is  not  closed  in  until  the  last  moment. 

The  total  number  of  tuyere  openings  is  14 — 6  behind,  4  in 
front,  and  2  at  each  end.  The  interior  rectangle  is  3  feet  5  inches 
wide  and  11  feet  8  inches  long,  although  any  exact  adherence  to 
these  measurements  is  unnecessary,   the   interior  of   the  furnace 

Fig.  46.— Plan. 
being  soon  burned  out  into  an  irregular  shape  and  usually  much 
larger  than  the  size  just  given. 

Strong  tie-rods,  provided  at  their  extremities  with  loops,  and 
buried  deeply  in  the  foundation,  are  placed  in  position  as  indicated 
in  the  cut.  Unless  the  transverse  rods  can  be  placed  at  a  depth 
of  two  or  three  feet  below  the  surface,  they  should  merely  be  fas- 
tened into  the  wall  by  hooks,  as  they  would  certainly  be  melted 
away  in  time. 

The  brick  should  be  laid  with  the  closest  possible  joints,  and  in 
a  very  thin  mortar  made  of  half  each  of  raw  and  burned  fire-clay, 
ground  exceedingly  fine. 

Heavy  railroad  iron  may  be  used  for  binders,  and  should  be  used 
rather  more  than  less  liberally  than  shown  in  the  illustration,  as 

SCALE  fe  IN.  TO  THt'  FOOT 

Fig.  47. — The  Orfokd  Brick  Furnace. 



the  expansive  force  is  enormons  when  the  furnace  is  in  full  heat, 
and  any  serious  cracking  tends  greatly  to  shorten  its  existence. 

If  tire-brick  are  expensive,  the  outside  lining,  above  the  panels, 
and  to  a  depth  of  12  inches,  may  be  constructed  of  red  brick, 
although  this  is  not  recommended. 

The  usual  height  from  the  tuyeres  to  the  threshold  of  the 
charging-door  is  8  feet;  but  this,  of  course,  may  be  varied  to  suit 
the  character  of  the  ore  to  be  smelted.  The  charging-doors  are 
three  in  number  and  of  large  size.  All  further  details  of  construc- 
tion are  plainly  shown  in  the  cut. 




i ^^^^^^^^^^^^H 


ll  II    ii 








Fig.  -48. — The  Orford 


Raschette"  Fcrnace. 

The  chimney  should  never  be  made  smaller  than  here  shown, 
and  if  a  vertical  down-take  is  used,  connected  with  flues  for  the 
saving  of  the  tiue-dust,  its  dimensions  should  be  increased  one- 
third.  The  latter  construction  is  much  preferable  to  the  simple 
vertical  chimney,  and  is  absolutely  essential  where  anything  but 
the  poorest  material  is  smelted,  as  the  loss  in  flue-dust,  owing  to 
the  enormous  volume  of  blast  peculiar  to  this  practice,  is  very 
great — especially  as  a  largo  proportion  of  the  charge  often  consists 
of  fine  ore,  it  having  been  found  that  these  large  rectangular  fur- 
naces are  peculiarly  adapted  to  the  treatment  of  that  material. 

The  tnyeres  consist  of  rather  heavy,  galvanized  sheet-iron — No. 
18 — and  are  connected  with  the  vertical  branches  of   the  main 



blast-pipe  surrounding  the  furnace,  with  thick  duck  tuyere-bags, 
soaked  in  a  strong  solution  of  alum  to  render  them  less  inflammable 
and  to  fill  the  pores  of  the  cloth.  Their  diameter  may  vary  with 
the  character  of  the  ore  under  treatment,  but  is  usually  from  five 
to  six  inches,  the  pipes  being  merely  thrust  a  short  distance  into 
the  square  orifices  left  in  the  brick-work,  and  made  tight  with 
plastic  clay. 

There  remains  nothing  in  the  construction  of  this  furnace  that 
cannot  be  plainly  seen  from  the  illustrations,  and  the  discussion  of 





fr        't\    f^       Fl    fl       R 











ScaJLe  %  in  tp  the  foot 

Fig.  49. 

its  management  from  the  time  when  taken  in  hand  by  the  smelter 
will  now  be  proceeded  with. 

It  is  frequently  customary  to  form  the  bottom  of  a  solid  mass  of 
fire-brick,  placed  on  end,  and  brought  up  to  within  10  inches  of 
the  tuyere  openings,  sloping  slightly  toward  the  slag-run  in  the 
center  of  the  front  wall.* 

*The  practice  of  basing  the  bottom  upon  an  arcli  built  over  an  open  space 
below  must  be  strongly  condemned,  as  it  will  simply  result  in  the  cutting 
through  of  the  arch,  and  the  total  disappearance  of  all  metal  until  the  cavity  is 
filled,  making  eventually  a  solid,  but  somewhat  expensive  bottom. 


The  author  has  found  the  following  method,  practised  originally 
by  the  Orford  Company,  far  superior  to  any  other,  especially  where 
low-grade  matte  is  to  be  produced,  the  most  difficult  of  all  copper- 
bearing  materials  to  confine  within  brick  walls. 

After  filling  in  the  foundation  with  betou  to  a  foot  below  the 
ground  level,  the  furnace  bottom  is  begun  by  laying  two  courses  of 
fire-brick  on  end,  and  with  the  closest  possible  joints.  This  still 
leaves  a  space  of  from  'Z-i  inches  to  30  inches  to  bring  the  bottom 
to  the  proper  height,  which  is  filled  in  as  follows: 

The  furnace  and  foundation  being  thoroughly  dried  by  at  least 
four  days'  brisk  firing  with  brands  and  similar  material,  enough 
coke  is  dumped  into  the  red-hot  shaft  to  fill  it  to  a  point  some 
three  feet  above  certain  temporary  openings  that  should  be  left  in 
the  brick-work  while  building.  These  openings  correspond  in 
size,  number,  and  position  with  the  permanent  tuyere  openings, 
except  that  they  are  some  8  inches  lower  and  directly  beneath  the 
regular  orifices,  which,  for  the  present,  are  plugged  with  clay. 

Some  six  or  eight  tons  of  calcined  quartz  crushed  to  the  size  of 
chestnuts  and  mixed  witii  about  5  per  cent,  of  fusible  slag,  are 
spread  upon  the  coke;  and  as  soon  as  the  latter  is  properly  on  fire 
above  the  temporary  tuyere  openings,  the  blast-pipes  are  put  in 
place,  and  a  light  blast  is  continued  until  the  coke  is  burned  away, 
and  the  sticky,  half-melted  charge  threatens  to  flow  into  the  tuyere 
openings.  The  unconsumed  coke  and  excess  of  quartz  are  removed 
through  the  breast  panel — which  was  built  up  temporarily  of 
4-inch  brick-work;  and  the  furnace,  being  tightly  closed,  is  al- 
lowed to  cool  very  gradually  for  twenty-four  hours  or  more. 

If  the  operation  is  successful,  the  bottom  will  be  as  solid  and 
infusible  as  can  be  made,  nor  will  any  attempt  at  the  substitution 
of  basic  material  for  quartz,  in  consideration  of  the  probably 
highly  ferruginous  character  of  the  slag  to  be  produced,  result  in 
any  improvement  on  the  plan  recommended. 

It  is  probably  as  good  a  bottom  as  can  be  made,  although,  as 
will  be  later  seen,  it  offers  but  little  resistance  to  a  hot  low-grade 
matte,  when  produced  at  the  rate  of  from  30  to  50  tons  daily. 

The  furnace  being  thoroughly  dried  and  heated,  blowing  in 
may  follow  at  once,  it  being  only  necessary  to  plug  the  temporary 
tuyere  orifices,  fill  the  shaft  with  coke  to  a  point  some  3  feet  above 
the  permanent  tuyeres,  and  allow  the  fire  to  ascend  to  these  open- 
ings before  filling  the  shaft  with  alternate  layers  of  charge  and 
fuel,  and  putting  on  a  light  blast.     All  this  may  be  done  the  night 


before  startiug,  and  the  forehearth,  with   siphou-tap,  must   then 
be  arranged.     (See  section  on  "Forehearths.") 

The  full    burden  may   be   reached    after  feeding    two    quarter 
charges,  four  half  charges,  and  eight  three-quarter  charges,  slag 
being  substituted  for  ore  to  a  considerable  extent,  until  the  condi- 
tion of  the  furnace  warrants  the  employment  of  the  normal  mixture. 

This  is  shown  by  the  gradual  change  of  the  color  of  the  slag 
from  a  dull  red  to  a  yellowish  white;  the  entire  ceasing  or  great 
diminution  of  smoke  arising  from  the  slag;  a  certain  peculiar  vis- 
cosity (except  in  very  basic  slags)  when  it  falls  into  the  pot;  a 
general  brightening  of  the  tuyeres,  succeeded  by  the  formation  of 
short  noses,  perforated  abundantly  with  bright  holes;  and  a  steady 
and  rapid  sinking  of  the  charge. 

Although  the  charging  of  the  blast-furnace  is  always  one  of  the 
most  important  manipulations  belonging  to  this  apparatus,  it  is 
doubly  the  case  with  the  furnaces  now  under  discussion. 

While  the  walls  of  the  water-jacket  are  thoroughly  protected 
and  entirely  unassailable,  the  mason-work  of  the  brick  furnace  is 
completely  exposed,  and  any  error  in  the  proportion  of  fuel  to  ore, 
or  in  the  manner  of  charging,  is  sure  to  be  followed  by  serious 

This  is,  strange  as  it  may  seem,  peculiarly  tlie  case  with  a  siliceous 
charge,  and  nothing  can  more  clearly  illustrate  the  proper  method 
of  working  than  a  brief  description  of  an  irregularity  that  is  con- 
stantly liable  to  occur,  and  that  will  be  quickly  recogrized  by  all 
practical  cupola  smelters. 

An  imaginary  case  will  be  assumed  where  a  newly  blown-in  fur- 
nace in  good  condition,  but  with  a  slightly  too  siliceous  charge, 
begins  to  become  too  hot  in  one  end,  through  some  slight  irregu- 
larity of  feeding,  or  through  an  improper  proportion  of  ore  to  fuel 
— either  too  much  or  too  little  of  the  same  ijrodncing  very  similar 

The  attention  of  the  foreman  will  be  called  to  the  fact  that  one 
of  the  end  panels  is  becoming  very  hot,  which,  as  it  consisis  of  18 
inches  of  fire-brick,  shows  either  that  the  inner  temperature  is 
much  too  high  or  that  the  bricks  have  already  been  thinned  by 

A  glance  into  the  tuyere  opening  shows  that  a  heavy  black  nose 
has  already  formed,  resulting;  from  the  fusion  of  the  fire-brick 
above,  which  form  a  crust  almost  impervious  to  a  steel  bar,  and 
exceedingly  infusible. 


A  consultation  with  the  man  who  feeds  that  end  of  tlie  furnace 
will  elicit  the  information  that  that  portion  of  the  charge  is  sink- 
ing very  slowly,  and  tliat  the  heat  is  rising  to  the  surface. 

At  the  same  time,  the  blast-gauge  will  show  an  increased  ten- 
sion, owing  to  the  blocking  up  of  the  tuyeres  that  supply  that 
portion  of  the  apparatus,  and  the  agglomeration  of  the  charge 
above,  owing  to  the  rapidly  ascending  temperature. 

The  already  too  siliceous  slag  is  rendered  still  more  infusible  by 
tlie  adinixtnre  of  silicate  of  alumina  from  the  melting  fire-brick; 
and  the  high  temperature  and  powerful  reducing  atmosphere,  re- 
sulting from  the  almost  stationary  condition  of  this  portion  of  the 
charge,  soon  begin  to  reduce  metallic  iron  out  of  the  slag,  and 
even  from  the  matte,  the  sulphur  being  driven  away  to  a  consider- 
able extent  by  the  powerful  blast,  high  temperature,  and  slow 
removal  of  the  molten  products. 

The  slimy,  half-fused  metallic  iron  is  soon  recognized  by  the  bar 
which  is  constantly  thrust  into  the  choked  tuyeres,  and  the  inex- 
perienced metallurgist,  following  the  teaching  of  all  our  best  text- 
books, reasons  that  the  reduction  of  iron  comes  from  too  highly 
ferruginous  a  charge,  and  destroys  all  hope  of  improvement  by 
cutting  off  a  portion  of  the  iron  from  the  charge  fed  into  that  end 
of  the  furnace. 

This  further  diminution  of  the  oxide  of  iron,  and  consequent 
necessary  increase  of  temperature  to  rnelt  the  more  and  more  infu- 
sible slag,  soon  bring  about  the  exact  conditions  prevailing  in  an 
iron-ore  blast-furnace.  Metallic  iron  is  reduced  in  large  quantities, 
while  the  temperature  is  raised  several  hundred  degrees,  before 
the  slag — now  virtually  an  acid  silicate  of  alumina  and  lime — will 
become  sufficiently  softened  to  run  at  all.  In  the  meantime,  the 
furnace  wall,  at  the  panel,  is  burned  nearly  through;  Jets  of  blue 
flame  appear  at  every  joint  and  crevice,  and  the  most  superficial 
examination  shows  that  the  process  is  extending  into  one  or  the 
other  of  the  corner  columns,  threatening  the  stability  of  the  struc- 
ture, and  still  more  alarming  the  person  in  charge.  The  column 
of  ore  in  that  end  of  the  furnace  hardly  sinks  at  all;  the  heat  is 
ascending  to  the  surface  of  the  charge;  and  the  general  increased 
stickiness  of  the  rapidly  lessening  slag-stream,  increase  in  tenor  of 
the  matte,  and  deposition  of  lumps  of  metallic  iron  in  one  or  both 
compartments  of  the  forehearth,  show  that  the  end  is  not  far  off 
and  unfold  the  near  prospect  of  a  chilled  furnace,  and  the  probable 
presence  of  a  block  of  half-molten  ore  and  iron  that  is  almost  im- 


pervious  to  tools,  aud  may  result  in  the  entire  abandonment  and 
destruction  of  the  furnace. 

This  is  one  of  the  most  common  aud  well-known  occurrences  in 
small  furnaces  and  with  inexperienced  metallurgists,  and  might 
just  as  well  happen  to  the  large  furnaces  now  under  discussion, 
were  it  not,  fortunately,  that  their  construction  and  management 
are  not  likely  to  be  undertaken  except  by  men  of  experience,  and 
also  that,  owing  to  their  greater  size,  a  threatening — or  even  estab- 
lished— chill  is  much  more  easily  managed  than  in  the  case  of  the 
smaller  cupolas,  whose  contracted  shaft  is  filled  up  solid  almost 
before  one  is  aware  that  anything  is  going  wrong. 

Owing  to  the  great  area  of  the  Orford  furnace,  a  considerable 
portion  of  the  shaft  may  be  completely  blocked  by  a  chill,  while 
a  brisk  fusion  is  progressing  in  the  other  half,  giving  an  opportu- 
nity, by  the  use  of  skill  and  experience,  to  gradually  smelt  away 
the  solidified  portion  and  eventually  bring  matters  back  to  their 
normal  condition. 

Returning  to  the  imaginary  case  that  has  just  been  followed  to 
a  disastrous  termination,  the  writer  will  endeavor  to  show  how 
such  a  catastrophe  may  be  averted,  and  will  describe  the  course  of 
events  as  they  have  occurred  scores  of  times  to  every  practical 

The  moment  that  it  is  noticed  that  one  end  or  corner  of  the 
furnace  is  becoming  abnormally  hot,  and  that  the  column  of  ore 
corresponding  thereto  is  sinking  slowly,  the  tuyeres  belonging  to 
that  portion  of  the  shaft — from  one  to  three  in  number — are  im- 
mediately removed,  and  the  openings  slightly  plugged  with  clay. 
At  the  same  time,  several  charges  of  the  most  fusible  slag — that 
from  matte  concentration  and  containing  a  very  high  perceritage 
of  iron  is  best — are  given,  in  place  of  ore,  and  the  whole  furnace 
is  most  carefully  Avatched,  to  learn  whether  the  burning  is  due 
merely  to  some  local  irregularity  in  feeding,  or  whether  some  im- 
portant point  affecting  the  whole  process  is  at  fault;  such  as  too 
much  or  too  little  fuel  in  proportion  to  ore;  improper  composition 
of  slag;  incorrect  feeding;  too  stronger  too  weak  a  blast,  etc.,  etc. 

Experience  alone  can  qualify  the  metallurgist  to  quickly  and 
correctly  detect  the  cause  of  the  trouble  and  apply  the  appropriate 
remedy;  but  in  any  case,  if,  after  taking  the  precautions  enumer- 
ated and  waiting  a  sufficient  time  to  get  their  full  effect,  the  burn- 
ing still  continues,  it  becomes  evident  that  the  trouble  is  deep- 
seated  and  of  some  extent. 


Vigorous  measures  are  therefore  required  to  stop  the  melting  of 
the  brick-work  above  the  tuyeres,  and  not  only  to  cool  down  the 
heated  end  of  the  furnace,  but  also  to  repair,  as  far  as  possible, 
the  damage  already  done  to  the  pauels — or  even  to  the  corners  of 
the  main  columns. 

Still  keejiing  the  offending  tuyeres  closed  as  already  described, 
a  full  charge  of  siliceous  ore  should  be  fed  in  such  a  way  that  it 
will  sink  to  the  indicated  spot.  This  may  be  given  either  with  or 
without  coke,  or  may  be  followed  by  a  second  or  third,  or  even  a 
greater  amount,  as  the  crcumstances  indicate;  proceeding  with 
extreme  caution,  and  allowing  some  two  hours  to  intervene  between 

The  author  has  found  it  necessary  to  charge  as  much  as  11  tons 
of  almost  pure  silica — quartz  with  specks  and  veiuletsof  carbonates 
and  oxides  of  copper — into  one  corner  of  an  overheated  furnace, 
and  this  entirely  without  coke,  before  the  gradual  cooling  of  the 
external  walls,  normal  and  even  sinking  of  the  charge,  and  lower- 
ing of  the  temperature  at  the  charging-door,  indicated  that  the 
mischief  had  ceased. 

The  office  of  this  siliceous  addition  is  not  to  render  the  slag  in 
general  more  siliceous.  This  would  only  bring  about  the  evils 
already  indicated,  and  probably  cause  a  heavy  reduction  of  metallic 
iron.  Its  object  is  rather  to  produce,  by  the  sudden  arrival  of 
such  a  body  of  cold,  infusible  material,  such  an  overwhelming  effect 
as  completely  to  cool  down  that  portion  of  the  shaft,  the  silica 
itself  softening  somewhat  and  remaining  for  the  most  part  in  the 
corner  of  the  furnace  corresponding  to  the  point  over  which  it  was 
cnargud.  It  attaches  itself  to  the  walls  and  bottom,  and  fills  up 
the  cavity  caused  by  the  fusion  of  the  fire-brick,  lowering  the  tem- 
perature at  the  same  time  to  a  considerable  extent,  but  producing 
no  marked  effect  on  the  general  character  of  the  slag. 

When  this  operation  is  successful,  as  is  usually  the  case,  the 
thinned  and  heated  brick-work  is  virtually  restored,  the  deeply 
excavated  bottom  is  filled  up  to  the  general  level,  and  matters  re- 
sume their  normal  condition,  all  irregular  bunches  and  protuber- 
ances of  the  siliceous  addition  that  may  have  adhered  to  the  fur- 
nace walls  becoming  gradually  melted  away  and  smoothed  down, 
until  the  interior  mason-work,  if  visible,  would  be  seen  to  have 
almost  assumed  its  original  appearance. 

Such  a  result  may  seem  very  doubtful,  and,  in  fact,  the  whole 
operation  may  appear  to  partake  too  much  of  the  marvelous  to 


those  unfamiliar  with  such  practice.  The  author  would  hesitate 
before  describing  the  foregoing  operation  as  a  matter  of  general 
everyday  occurrence,  were  it  not  that  it  can  be  vouched  for  in  its 
entirety  by  a  considerable  number  of  well-known  and  reliable  gen- 
tlemen. This  jiractice,  as  initiated  by  certain  members  of  the 
Orfcrd  Company,  already  mentioned,  has  spread  until  it  is  now  a 
well-known  and  recognized  part  of  our  local  copper  metallurgy. 
The  skill  attained  by  certain  foremen  in  managing  these  very 
large  furnaces  is  quite  remarkable,  and  far  beyond  anything  de- 
scribed in  this  treatise. 

While  the  imaginary  case  just  described  in  detail  represents  only 
one  of  the  various  accidents  peculiar  to  all  forms  of  blast-furnace, 
it  still  is  at  the  bottom  of  a  very  large  proportion  of  the  instances 
of  "freezing,"  "choking-up,"  "burning-ont,"  etc.,  etc.  Paradox- 
ical as  it  may  appear,  the  two  common  accidents  of  "burning-out" 
and  "freezing-up"  are  closely  connected,  and  in  reality  only  two 
different  stages  of  the  same  morbid  process.  The  young  metallur- 
gist cannot  overestimate  the  importance  of  the  fact  that  it  is  quartz 
in  one  or  another  of  its  forms,  in  a  furnace  that  is  not  intended 
for  a  siliceous  charge,  that  is  the  most  frequent  cause  of  smelting 
difficulties  and  disasters.  Seven  out  of  the  last  eight  cases  of 
metallurgical  difficulties  for  which  the  writer  was  called  upon  to 
prescribe,  were  due  to  this  cause. 

In  spite  of  the  frequency  and  apparent  simplicity  of  this  diffi- 
culty, some  smelters  of  experience  never  seem  to  have  learned  the 
cause,  and  attribute  the  slow  and  irregular  running  of  the  cupola 
and  the  frequent  filling  up  of  the  crucible  with  sows  to  "too  much 
iron  in  the  charge" — "too  much  suljihur" — •"magnesia  in  the  lime- 
stone flux,"  etc.,  when  in  almost  every  instance  a  mere  ocular 
examination  of  the  slag  is  sufficient  to  show  that  silica  is  at  the 
bottom  of  the  trouble.  No  apology  is  needed  for  emphasizing  this 
point  when  men  considered  as  expert  metallurgists  are  cojistantly 
falling  into  this  error. 

It  is  especially  during  such  accidents  and  irregularities  that  the 
great  advantages  of  these  very  large  furnaces  become  fully  appar- 
ent. Where  a  small  shaft  Avould  soon  be  completely  and  irretriev- 
ably choked,  necessitating  the  great  expense  of  blowing  down  and 
subsequently  chiseling  out  the  half-fused  mass  of  ore  and  cinder, 
no  large  furnace,  in  any  instance  known  to  the  author,  has  ever 
become  so  blocked  up  and  filled  with  a  chill  that  it  has  not  been 
quite  easy  to  save  it  by  using  appropriate  means.     Even  though 



oue  end  be  completely  blocked,  there  is  always  ample  spa«:e  at 
some  points  of  its  eleven-foot  shaft  to  permit  the  descent  of  the 
charge  and  retain  a  sufBcient  number  of  tayeres  intact  to  gradually 
melt  out  the  chill  and  restore  the  shaft  to  something  like  its  formci 
dimensions.  Some  considerable  irregularity  of  form  naturally  re- 
sults from  repeated  manipulations  of  this  kind;  but  so  long  as 
sufficient  area  remains  at  the  tuyere  level,  and  no  projecting  masses 
impede  the  regular  descent  of  the  charge,  no  diminution  of  capac- 
itv  need  follow,  nor  increase  of  difficulty  in  managing  the  furnace. 
The  accompanying  sketch  gives  a  tolerably  correct  view  of  tlie 
shape  of  one  of  these  large  brick  furnaces  at  the  tuyeres  upon  its 
blowing-out  for  repairs  after  a  continuous  campaign  of  8^  mouths, 
during  which  time  over  18,000  tons  of  exceedingly  ferruginous  ore 
were  smelted  in  it,  yielding  a  very  low-grade  matte  and  slag  aver- 

FiG.  50— The  Rectangle  Shows   the    Shape  Before  the  Campaign; 


aging  about  22  per  cent,  silica  and  over  70  per  cent,  protoxide  of 
iron.  As  it  is  drawn  to  a  scale,  the  extent  of  the  irregularity  is 
easily  appreciable,  the  original  dimensions  being  3  feet  3  inches  by 
11  feet  -4  inches. 

In  fact,  the  fall  capacity  of  this  type  of  furnace,  when  smelting 
a  basic  ore,  is  not  reached  until  the  walls  are  burned  ont  to  a  con- 
siderable extent,  which  may  indicate  the  policy  of  widening  the 
furnace  in  the  first  place.  When  smelting  a  siliceous  ore,  or  when 
a  large  proportion  of  fines  is  present,  the  gain  in  width  is  accom- 
panied with  a  decrease  of  temperature  and  irregularities  in  the 
descent  of  the  charge — circumstances  that  soon  rectify  the  trouble 
by  adhering  to  the  walls,  and  filling  up  the  shaft  again  with  a 
rapidity  that  may  be  disastrous  if  not  observed  and  remedied  in 

As  has  been  already  briefly  mentioned,  the  cutting  down  of  the 
bottom  and  piercing  of  the  foundation-walls  is  an  accident  that 


sometimes  occurs,  although  usually  only  wheu  the  charge  consists 
of  a  very  fusible  unroasted  ore,  producing  a  matte  of  low  grade — 
from  25  per  cent,  downward — whose  fiery  and  corrosive  qualities 
are  well-kuowu  to  all  furnace-men.  It  is  to  the  great  quantity,  as 
well  as  corrosive  quality,  of  this  substance,  and  this  usually  in 
connection  with  a  basic  slag,  that  this  destructive  process  is  due; 
and  in  spite  of  much  care  and  expense  bestowed  on  the  matter,  no 
material  has  yet  been  found  that  will  withstand  a  daily  production 
of  from  20  to  45  tons  of  this  intractable  product.  But  a  means 
of  lessening  its  destructive  action,  as  well  as  of  greatly  pro- 
longing the  life  of  the  entire  structure  and  rendering  its  manage- 
ment much  easier,  lias  been  discovered  and  quite  generally  adopted, 
being  first  brought  into  notice  by  Mr.  John  Thomson,  of  the 
Orford  Company.  It  consists  in  duplicating  the  furnace  plant 
and  running  each  individual  cupola  only  ten  or  twelve  hours  of 
the  twenty-four.  This  is  a  scheme  that  seldom  recommends  itself 
to  one  on  first  hearing,  but,  after  a  thorough  trial,  will  be  found 
to  possess  numerous  important  advantages,  while  its  only  drawback 
is  the  increased  first  cost  of  the  plant — a  trifling  consideration  in 
comparison  with  the  large  interests  usually  at  stake. 

A  mere  doubling  of  the  cupola  plant  is  sufficient  to  overcome 
the  difficulties  mentioned;  but  if  it  be  desired  to  reap  the  full  ad- 
vantages of  the  scheme,  a  corresponding  increase  should  be  made 
in  the  blast  apparatus.  This  being  effected,  the  entire  smelting 
process  may  be  confined  to  the  daytime,  avoiding  the  difficulties 
and  drawbacks  of  night  work,  saving  the  wages  of  one  or  more 
foremen,  and  rendering  it  possible  for  the  manager  to  retain  that 
complete  personal  oversight  of  the  smelting  process  that  is  unat- 
tainable when  half  of  it  is  concealed  from  his  inspection.  If  this 
were  the  only  benefit  derived  from  the  above  plan,  it  would  in 
most  cases  be  well  worthy  of  adoption;  but  the  advantages  accru- 
ing to  the  furnaces  themselves,  as  well  as  to  the  entire  process,  are 
too  numerous  and  far-reaching  to  be  thoroughly  explained  in  this 

In  the  first  place,  the  cutting  down  of  the  furnace  bottom  is 
usually  completely  remedied  by  the  long  and  ever-recurring  periods 
of  complete  repose,  during  which  the  thinned  brick- work  is  again 
sealed  by  the  chilling  of  the  molten  products;  the  hearth  is  re- 
newed by  the  solidification  of  the  matte  and  slag  still  remaining  in 
the  cavities  of  the  hearth;  the  overheated  brick-work  cools  from, 
the  outside  to  such  an  extent  that  the  area  that  to-day  has  given: 


-constant  annoyance  by  its  obstinate  burning,  with  the  constant 
threat  of  finally  breaking  through  and  causing  serious  trouble,  will 
to-morrow  be  found  as  cool  as,  or  cooler  than,  any  other  portion, 
owing  to  the  thinness  of  its  walls;  and  various  slight  difficulties 
that  are  pretty  sure  to  occur  in  the  course  of  a  long  run  are  averted 
before  they  become  of  importance,  while  the  trouble  begins  at  a 
new  point,  only  to  be  again  averted  before  it  has  gained  serious 
headway.  This  is  by  no  means  an  uncommon  or  imaginary  case, 
but  a  matter  of  frequent  occurrence,  and  these  lines  are  written 
after  several  years'  trial  of  both  the  constant  and  intermittent 
metiiod  of  smelting,  the  experience  of  others  who  have  fairly  tried 
this  plan,  in  connection  with  large  brick  furnaces,  being  equally 

The  writer's  attention  was  first  called  to  this  matter  in  1871, 
when  noticing  the  almost  invariable  improvement  in  behavior  and 
capacity  that  succeeded  any  accidental  stoppage  of  cupola-furnaces 
that  he  was  then  managing.  The  ores  were  exceedingly  bad  and 
siliceous,  and  the  difficulties  detailed  in  thepriceding  pages  fol- 
lowed each  other  with  disheartening  regularity  and  frequency. 
Great  pains  were  taken  to  secure  a  steady  and  uninterrupted  run, 
fears  being  entertained  that  any  stoppage  would  be  disastrous  to 
the  furnace  in  the  more  or  less  critical  condition  that  seemed  to  be 
its  normal  state;  but  after  finding  that  the  benefits  following  any 
temporary  stoppage  of  the  machinery  had  become  so  obvious  that 
the  foreman  was  in  the  habit  of  purposely  causing  slight  accidents 
in  order  to  help  his  furnace  out  of  some  particularly  critical  situa- 
tion, it  was  decided  to  adopt  the  practice  of  stopping  for  two  or 
three  hours  whenever  the  ordinary  incidents  of  burning  out,  etc., 
became  unusually  critical.  This  habit  was  carried  further  and 
further,  proceeding  with  caution  and  gradually  lengthening  the 
stoppages,  until  it  came  to  be  considered  an  almost  universal 
remedy,  and  was  as  often  applied  for  chilling  or  freezing  up  as  for 
the  opposite  condition  of  affairs,  and  no  misfortune  ever  arose  from 
its  reasonable  application. 

This  practice,  like  every  other,  must  be  used  with  care  and 
judgment,  and  may  easily  be  carried  to  an  extreme,  bat,  as  a  rule, 
is  the  least  dangerous  measure  that  can  be  adopted  with  a  badly 
acting  furnace  of  large  area.  A  small  furnace  might  easily  chill 
in  a  few  hours,  so  that  the  length  of  the  period  of  repose  must  be 
proportioned  to  the  size  of  the  shaft  and  to  the  cubic  contents  of 
•the  heated   material.     The  thickness  of  the  walls  must  also  be 


considered,  as  the  rapidity  of  the  escape  of  heat  depends  upon  the 
thickness  of  the  brick-work.  It  is  hardly  necessary  to  say  that 
every  orifice  and  crevice  about  the  furnace  must  be  tightly  sealed, 
the  tuyeres  being  removed,  and  their  openings,  as  well  as  the  slag- 
run,  being  tightly  filled  with  damp  clay,  while  the  brick-work  in 
their  vicinity  must  be  searched  for  possible  cracks,  and  all  such 
openings  carefully  plastered  over.  Otherwise,  the  incoming  cur- 
rents of  air  would  gradually  burn  away  all  the  fuel  contained  in 
the  charge,  leaving  the  furnace  in  a  hopeless  condition.  If  it  is 
to  stand  still  any  length  of  time,  such  as  over  night,  a  little  extra 
■coke  should  be  given  an  hour  or  two  before  stopping,  so  that  there 
may  be  an  abundance  of  fuel  in  the  bottom  of  the  furnace.  A 
small  charge  of  basic  slag  should  also  be  given;  and  as  soon  as  the 
blast  is  taken  off,  the  basin  or  forehearth  tapped,  and  all  openings 
sealed,  the  surface  of  the  charge  should  be  covered  with  a  layer  of 
fine  coke,  over  which  is  spread  an  inch  or  two  of  fine,  fusible  ore. 
The  slag-hole  connecting  the  furnace  with  the  forehearth  should 
be  thoroughly  cleared  out;  the  layers  of  chilled  slag  and  ashes,  by 
which  the  blowing  through  of  the  blast  is  prevented,  removed,  and 
the  channel  itself  filled  with  fine  charcoal  or  coke,  well  rammed  in 
with  a  "stopping  pole."  This  is  rendered  impervious  to  air  by  an 
-exterior  plug  of  clay,  and  the  forehearth,  while  still  hot,  being 
scraped  clean  of  all  half-fused  masses  of  slag  or  reduced  iron,  and 
everything  being  prepared  for  the  morrow's  work,  the  cupola  may 
be  left  in  charge  of  an  experienced  watchman — preferably  an  old 
smelter.  On  the  ensuing  morning,  a  light  blast  is  put  on,  and  the 
channel  being  cleared  out,  slag  will  flow  in  from  five  to  ten  min- 
utes, while  in  half  an  hour  the  furnace  will  be  in  normal  condition, 
and  in  most  cases  smelting  more  rapidly  and  satisfactorily  than 
when  left  the  previous  evening. 

The  extreme  length  of  time  that  a  large  furnace  may  stand  in 
this  way  without  injury  is  unknown  to  the  author.  Much  de- 
pends on  the  fusibility  of  the  charge,  the  character  of  the  fuel, 
the  moreor  less  perfect  exclusion  of  all  air,  and  probably  also  upon 
the  quality  and  amount  of  sulphide  compounds  present,  whose 
gradual  oxidation  may  sustain  the  vitality  of  the  charge  for  a  much 
greater  length  of  time  than  if  absent.  The  following  instances, 
from  personal  experience,  show  that  a  considerable  delay  is  per- 

A  furnace  running  on  a  fusible  charge  of  calcined  pyritic  ore 
■was  shut  down  Friday  noon,  on  account  of  an  accident  to  the 


engine.  Farther  examination  showed  the  accident  to  be  of  such 
a  nature  as  to  cause  a  delay  until  the  succeeding  Wednesday  night 
— o^  days — at  the  end  of  which  time  a  light  blast  was  applied 
without  much  hope  of  a  favorable  result,  although  the  coke  on  top 
of  the  charge  was  hot  and  glowing. 

There  seemed  a  good  deal  of  obstruction  to  the  blast  at  first; 
but  in  twenty  minutes,  a  cold,  thick  slag  began  to  run,  which 
gradually  improved,  until  the  furnace  resumed  its  normal  condi- 
tion and  capacity  in  about  eight  hours.  The  charge  had  sunk 
about  two  feet  in  the  furnace  during  this  period  of  repose.  The 
grade  of  the  tirst  tap  of  matte  (the  siphou-tap  being  impracticable 
in  this  condition  of  affairs)  was  46  per  cent.,  the  ordinary  average 
being  from  28  to  29  per  cent.  The  succeeding  tappings  gradually 
decreased — going  successively  42,  37,  and  34  per  cent.,  the  normal 
grade  being  reached  soon  after  the  furnace  had  regained  its  usual 

Periods  of  4  days,  3|,  3^,  3,  and  of  less  time,  appear  in  the 
writer's  notes,  the  only  serious  accident,  occurring  during  one  of 
the  shorter  periods,  being  caused  by  the  falling  out  of  two  of  the 
tuyere-plugs,  whereby  a  current  of  air  entered  the  furnace  for 
twelve  hours  before  being  discovered.  The  coke  was  completely 
burned  out  of  the  lower  portion  of  the  charge  for  about  two-thirds 
of  that  part  of  the  shaft  nearest  the  opening;  but  the  furnace  was 
eventually  saved  by  blowing  lightly  into  three  tuyeres  at  the  oppo- 
site end,  which  were  still  supplied  with  fuel,  and  little  by  little 
smelting  out  the  entire  half-fused  block  of  charge.  Much  benefit 
was  derived  by  introducing  coke  into  the  furnace  through  such 
tuyeres  as  seemed  to  warrant  the  trouble.  Owing  to  the  great  size 
of  the  tuyere  openings  [6  inches),  this  was  easily  effected,  and  the 
smelting  much  facilitated.  In  fact,  if  any  cavity  in  the  semi-fused 
mass  could  have  been  found  at  any  point  accessible  to  the  blast, 
nothing  would  have  been  simpler  than  to  break  a  hole  through 
one  of  the  brick  panels  and  fill  the  opening  with  coke.  The  author 
has  done  this  in  later  instances  with  very  satisfactory  results,  a 
cavity  opposite  the  tuyeres  having  been  formed  by  dragging  out 
a  lot  of  the  stock,  from  which  the  coke  had  burned  so  gradually  as 
not  to  fuse  it. 

Space  is  wanting  for  a  description  of  the  use  of  petroleum,  gas, 
and  other  concentrated  fuels  for  similar  purposes,  as  the  writer's 
own  experience  with  such   measures  has  been  entirely  unsatisfac- 


tory,  nor  can  he  find  any  record  of  successful  cases  in  the  annals 
of  American  copper  smelting. 

The  most  herculean  efforts  are  warrantable  when  any  reasonable 
probability  exists  of  the  saving  of  an  iron  furnace  from  complete 
chilling  up;  but  in  copper  smelting,  the  comparative  cheapness 
and  simplicity  of  the  structure  itself,  and  the  certainty  of  being 
able  to  remove  the  worst  chill  by  mechanical  means  in  a  compara- 
tively short  time,  render  such  unusual  and  expensive  measures  less 

The  oxidation  of  the  sulphides  in  the  charge  during  the  period 
of  repose  is  an  element  of  some  importance,  although  seldom  so 
striking  as  in  the  case  just  mentioned.  Still,  the  closing  down  of 
the  cupola  over  night  is  invariably  accompanied  with  a  perceptible 
rise  in  the  grade  of  the  matte  produced  during  a  certain  period 
succeeding;  being  greatest  at  first,  and  gradually  diminishing  as 
the  contents  of  the  furnace  are  replaced  with  fresh  ore.  This 
increase  in  richness  is  at  first  seldom  less  than  5  per  cent.,  dimin- 
ishing rapidly,  however,  as  the  ore  nearest  the  bottom  of  the  charge 
has  experienced  the  most  thorough  oxidation. 

Though  apparently  a  trivial  matter,  this  enrichment  of  the 
matte  is  a  direct  pecuniary  gain,  and,  according  to  a  rough  esti- 
mate, will  offset  the  interest  on  the  capital  necessary  for  the  double 
plant  several  times  over  in  the  course  of  a  year. 

Another  useful  and  frequently  applied  remedy  for  various  irreg- 
ularities in  cupola  smelting  is  the  so-called  "running  down"  of 
the  furnace,  by  which  is  meant  a  mere  cessation  of  charging  until 
the  column  of  ore  and  fuel  has  sunk  to  a  point  far  below  its  nor- 
mal limits.  The  shaft  is  then  rapidly  filled  with  the  usual  alter- 
nate charges  of  ore  and  fuel,  and  everything  goes  on  as  before. 

This  practice  is  sometimes  of  great  advantage,  obstinate  irregu- 
larities often  being  conquered  thereby,  and  the  normal  condition 
of  things  resumed.  It  is  especially  useful  when  it  is  desired  to 
create  a  sudden  and  profound  lowering  of  temperature  at  some 
point  where  a  serious  localized  burning  is  taking  place;  for  the 
exposure  of  the  naked  inclosing  walls  of  the  shaft  renders  it  possi- 
ble to  deposit  the  batch  of  ore  that  is  used  to  cool  the  walls  in  the 
exact  spot  where  it  is  needed;  and  it  is  possible  to  use  for  this 
purpose,  under  such  circumstances,  an  easily  fusible  ore  or  slag, 
instead  of  the  highly  siliceous  material  that  is  usually  selected 
when  this  process  of  cooling  down  is  undertaken  blindly  from 


Wall  accretions  may  also  be  reached  in  this  manner,  the  charge 
being  allowed  to  settle  until  they  are  exposed,  whereupon  they 
may  be  removed  by  a  long, bent  steel  bar  introduced  through  one 
of  the  charging-doors,  the  glowing  interior  being  cooled  down,  if 
necessary,  by  sprinkling  with  water. 

Still  another  means  of  remedying  the  cutting-down  of  the  fur- 
nace bottojn  has  been  mentioned  in  a  former  section,  but  is  some- 
times useful  in  connection  with  the  large  brick  furnace.  This  is 
the  iiitroduction,  through  the  tuyere  openings,  of  ore  or  sand, 
which,  being  both  cold  and  the  latter  infusible,  will  not  combine 
with  the  slag,  as  it  is  already  below  the  smelting  zone;  but  will 
simply  remain  in  place  and  assist  in  building  up  a  new  bottom. 
By  this  means,  even  the  molten  masses  present  may  be  partially 
solidified  and  a  great  advantage  gained  in  a  short  time.  The 
author  has  occasionally  tried  the  introduction  of  water  in  the  same 
manner  and  for  the  same  purpose,  taking  as  a  guide  the  very  de- 
cided local  chilling  produced  by  a  leaky  water-jacket;  but  the 
results,  though  locally  satisfactory,  are  not  sufficiently  extended, 
while  the  operation  itself,  especially  in  connection  with  a  low-grade 
copper  matte,  cannot  be  recommended  to  any  who  object  to  certain 
and  frequent  explosions  of  considerable  force. 

In  connection  with  the  measures  already  detailed  for  keeping 
the  furnace  in  proper  condition,  may  be  mentioned  the  external 
repairs  that  it  is  feasible  to  execute  while  the  furnace  is  still  in 
blast.  Not  all  smelters  are  aware  of  the  very  extensive  repairs 
that  may  be  carried  out  without  stopping  the  blast  more  than  a 
few  hours;  the  length  of  the  campaign  often  being  doubled  by  the 
construction  of  a  new  panel,  the  repairing  of  a  pillar,  and  other 
familiar  and  inexpensive  operations.  These  are  of  too  extensive 
and  varied  a  nature  to  be  enumerated  in  detail;  but  a  few  of  the 
te-ichings  of  experience  will  throw  some  light  on  the  practice  in 

The  replacement  of  one  or  more  panels  that  have  become  so 
thin  as  to  threaten  a  constant  breaking  through  of  the  charge  is  a 
simple,  though  very  hot  and  laborious  task. 

All  needful  material  for  the  renewal  being  prepared  and  collected 
on  one  spot,  the  blast  is  shut  off,  the  forehearth  tapped,  and  the 
condemned  brick-work  at  once  broken  in  with  sledge  and  bar.  So 
much  of  the  glowing  charge  as  is  necessary  is  at  once  dragged  out 
of  the  opening  with  long  hoes  and  rakes,  and  sprinkled  with  water 
so  that  the  men  can  stand  on  it  to  work. 


Wlien  the  bricks  have  been  removed  to  the  extent  deemed  neces- 
sary, the  cavity  left  in  the  column  of  stock  is  quickly  filled  with 
dampened  coke,  a  few  wooden  slats  being  wedged  across  the  open- 
ing, to  keep  the  fuel  from  falling  out. 

I'he  most  important  measure  is  to  obtain  a  solid  foundation  for 
the  new  wall,  and  to  acconiplish  this,  all  accretions  of  slag  and 
metal,  of  which  the  old  wall  largely  consisted,  must  be  chiseled 
away  until  sound  brick-work  is  reached,  which  being  leveled  with 
thick  tire-clay,  offers  a  proper  starting-point.  The  work  must 
proceed  with  great  rapidity,  as  the  passage  of  air  through  the 
opening  will  soon  consume  the  fuel  in  the  charge.  Little  atten- 
tion is  paid  to  neatness,  or  even  regularity,  so  long  as  strength 
and  tightness  are  obtained.  If  the  work  promises  to  occupy  more 
than  two  or  three  hours,  the  opening^hould  be  closed  at  the  begin- 
ning by  a  thin  plate  of  sheet  iron  tightly  cemented  at  the  edges 
with  clay,  outside  of  whicii  the  new  wall  is  raised.  When  all  is 
completed,  the  sheet-iron — unless  already  consumed — is  cut  away 
opposite  the  tuyere  openings,  and  the  blast  is  put  on  at  once,  there 
being  no  necessity  of  waiting  for  the  work  to  dry,  as  the  heat  from 
the  furnace  will  evaporate  all  moisture  quite  as  soon  as  is 

By  this  means,  extensive  repairs  may  be  executed  on  any  portion 
of  the  furnace,  it  being  even  possible  to  put  in  a  new  bottom,  or 
repair  the  foundation  walls,  by  suspending  the  charge  on  bars 
driven  transversely  through  the  furnace.  When  possible,  the 
ashes  of  the  rapidly  consumed  fuel  should  be  cleared  out  before 
starting  again;  but  there  are  but  few  instances  where  it  will  not 
be  found  better  to  blow  out  the  furnace  when  such  radical  repairs 
are  required. 

The  final  blowing  out  of  the  large  furnace  presents  no  peculiar 
features.  The  blast  should  be  lessened  as  the  charge  sinks,  and 
as  soon  as  slag  stops  running,  the  breast-wall,  and,  if  expensive 
repairs  are  imminent,  some  of  the  rear  and  end  panels  should  be 
knocked  in,  and  all  stock  and  fuel  dragged  out,  until  a  tolerably 
even  bottom  is  reached,  which  needs  no  preparation  for  the  suc- 
ceeding campaign. 

Any  burning  out  of  the  brick  pillars  that  form  the  main  sup- 
port of  this  furnace  should  be  carefully  watched  and  repaired 
before  it  has  proceeded  to  a  dangerous  extent.  This  burning  is 
sometimes  so  obstinate  that  when  it  is  important  not  to  sto})  the 
furnace  or  blow  out,  it  is  necessary  to  support  the  superincumbent 


brick-work  with  props  aud  braces,  which  should  reniaiu  in  place 
until  the  pillars  have  been  restored  to  their  former  strength. 

Estimates  of  the  cost  of  building  one  of  these  large  brick  fur- 
naces of  the  Orford  type  will  be  found  in  this  chapter. 

There  remains  to  be  still  considered  the  application  of  water 
tuyeres  and  other  cooling  devices  to  furnaces  constructed  of  brick 
or  stone. 

The  author's  own  experience  is  entirely  in  favor  of  the  employ- 
ment of  jJroperly  constructed  iron,  or  better,  broDze  or  copper 
tuyeres,  containing  a  space  for  the  introduction  of  water.  In  Col- 
orado and  other  places,  he  has  used  water  tuyeres  with  invariable 
satisfaction,  the  only  drawback  being  the  frequent  cracking  of  the 
cast-iron,  which  is  now  overcome. 

While  they  ofier  little  or  to  the  furnace  wall,  they 
are  indestructible  themselves,  and  by  delivering  the  wind  at  a 
fixed  point,  even  though  the  walls  may  be  eaten  away  all  about 
them  to  the  depth  of  a  foot  or  more,  they  remove  the  point  of 
greatest  heat  from  the  wall  itself,  and  practically  retain  the  smelt- 
ing area  at  the  same  invariable  size,  the  latter  being  practically 
bounded  by  vertical  planes  passing  through  the  nozzles  of  tlie 

It  is  also  possible,  if  desirable,  to  project  them  into  the  interior 
of  the  furnace  to  a  distance  of  several  inches  from  the  walls.  Al- 
though this  practically  diminishes  the  size  of  the  smelting  area,  it 
saves  the  walls  from  burning,  and  in  case  of  a  weak  blast  or  of  an 
unusually  dense  charge  arising  from  a  large  proportion  of  fine  ore, 
may  render  practicable  the  smelting  of  material  that  would  be 
impossible  under  other  circumstances. 

They  were  tried  on  the  first  large  Orford  furnaces,  but  failed, 
owing  to  the  severity  of  the  winter  and  other  accidental  causes, 
rather  than  from  any  fault  due  to  the  tuyeres  themselves.  Their 
construction  and  management  are  too  familiar  to  require  further 
explanation  in  these  pages. 

The  surface  cooling  of  the  brick-work  by  means  of  a  spray  of 
water  on  the  outside  has  been  tried  on  many  occasions  and  w'ith 
various  forms  of  apparatus.  It  has  rarely  given  satisfaction,  and, 
in  the  writer's  opinion,  is  as  dangerous  aud  worthless  a  device  as 
can  well  be  imagined.  To  those  familiar  with  the  results  of  con- 
tact between  water  and  molten  matte,  it  is  not  necessaiT  to  bring 
up  any  further  arguments  to  condemn  a  device  that  can  only  be 


accom|);mied  by  a  constant  wetting  of  everything  in  the  "vicinity 
o!  the  furnace. 

Besides,  the  idea  itself  is  an  extremely  faulty  one,  as,  owing  to 
tlie  non-conductivity  of  fire-brick,  a  wall  less  than  a  foot  thick 
may  continue  melting  on  one  side,  while  its  other  surface  is  con- 
stantly sjirayed  with  cold  water. 

All  devices  of  this  kind,  in  which  the  water  comes  in  contact 
with  the  free  exterior  sufrace  of  the  furnace  wall,  are,  in  the 
author's  opinion,  worse  than  useless,  and  likely  to  be  accompanied 
by  most  dangerous  results. 


Excavation  for  foundation:    1,000  cubic  feet  at  8  cents. . .  $80.00 

Foundation  of  beton 65.00 

Cubic  feet. 

Total  fire-brick  for  furnace  proper 1,640 

Lining  for  cross-flue  and  down-take     540 

Foreheartb,  etc 45 

Total    2,225 

At  18  brick  per  cubic  foot  =  40,050  at  $40  a  thousand  1,602.00 

Red  brick  for  down-take  and  flue:  16,800  at  $8 134.40 

6i  tons  fire-clay  at  $8 52.00 

6  casks  lime  at  $1.50 9.00 

2  tons  sand  at  $1.50 3.00 

Old  rails  for  binders:    180  yards  at  80  pounds  a  yard  == 

14,400  pounds  at  |  cent 108.00 

Tie  rods    for   furnace,   flue,  and  down-take:  620  Pounds. 

feet  of  li  iron  =  2,480  pounds 2,480 

Loops,  nuts,  etc 166 

Angle  iron  for  down-take 172 

Wrouglit-iron  rods,  etc.,  about  forelieartb 66 

Total 2,884 

At  2  cents  a  pound 57.68 

Castings:  Pounds. 

3  feed-door  frames 792 

Damper  and  frame 455 

Plates  for  foreheartb 560 

Slag  and  matte-spouts 80 

Plates  for  chargiug-floor 1,260 

Miscellaneous 420 

Total 3,567 

Brought  forward 2,111.08 


Carried  forward.  ? $2,111,08 

At  2^  cents  a  pound 89  17 

Material  and  labor  for  arch  patterns  and  other  carpenter 

work o  32.40 


Mason,  88  days  at  $4  ...   352.00 

Ordinary  labor,  102  days  at  $1.50  153.00 

9i  days,  smith  and  helper. 47.50 

Blast-pipe  and  tuyeres 136.00 

Cloth  for  tuyere  bags  and  labor 3.80 

Superintendence   120.00 

Miscellaneous 65.00 

Grand  total $3,109.95 

Tools  essential  to  furnace,  steel,  and  iron  bars,  shovels, 

rakes,  hamuaers 55.90 

15  slag-pots  at  $13.50 202.50 

4  iron  barrows  at  $9.00 36.00 

Manometer 2.50 

Total $296.90 

The  above  estimate  is  exclusive  of  main  blast-pipe,  blower, 
motive  power,  hoist,  and  chimney  or  dust-chambers;  tlie  allowance 
for  cross-floe  and  down-take  being  sufficient  to  cover  cost  of  chim- 
ney in  those  exceptional  cases  where  no  provision  is  made  for 
catching  the  immense  amonnt  of  flue-dust  generated  in  this 
method  of  smelting. 

A  compact  and  economical  hoist  and  ample  provision  for  a 
large  charging-floor  and  generous  bin  room  are  essential  to  con- 
venient and  economical  work. 



The  capacity  of  a  blast-fnrnace  is  dependent  upon  many  varying 
causes,  and  is  to  a  considerable  extent  independent  of  shape  or 
size,  though  its  tuyere  area  is,  of  course,  the  most  important 
factor  in  determining  the  amount  of  material  that  can  be  passed 
through  it. 

Next  to  the  fusibility  of  the  charge,  the  pressure  and  volume  of 
the  blast  have  the  principal  influence  in  determining  this  point, 
assuming  always  that  the  fuel  used  is  of  sufficient  strength  and 
density  to  permit  the  full  pressure  of  wind  that  may  be  found 
most  advantageous. 

Nothing  can  be  more  striking  than  the  change  in  the  rate  of 
smelting  of  a  large  cupola-furnace  as  the  wind  pressure  is  dimin- 
ished or  increased. 

The  author  has  taken  occasion  during  the  smelting  of  a  fusible 
charge,  and  with  the  furnace  in  perfect  condition,  to  ascertain  the 
difference  of  capacity  effected  by  changes  in  the  strength  of  the 

As  the  influence  of  the  change  is  almost  instantaneous,  it  is 
easy  to  arrive  at  such  flgures  with  considerable  accuracy,  measuring 
the  capacity  by  noting  the  number  of  pots  of  slag  produced  during 
periods  of  an  hoiu'  each,  and  with  varying  wind  pressure. 

The  following  table  shows  the  result  of  these  experiments  in  a 
compact  form,  repeated  sufficiently  often  under  varying  conditions 
to  establish  their  comparative  accuracy. 

It  should  be  mentioned  that,  in  order  to  insure  the  accuracy  of 
each  observation  independently  of  the  condition  of  the  furnace 
previous  to  the  experiment,  which  might  have  been  influenced  by 
the  preceding  test,  nearly  all  the  trials  were  made  at  different 
times,  but  with  the  furnace  as  nearly  at  its  normal  state  as  possi- 
ble, and  running  under  its  ordinary  pressure  of  blast-— about  10 
ounces  per  square  inch  : 






Blast  Pres-  Production  in 
sure  iuOz.          Tons. 
Per  Sq.  In.  Per  34  Hours. 

Assay  of  Slag 
in  Copper. 

Condition  of  Furnace  at  Close  of 


•s! ! ! ! 




~ — . 








Very  hot.    All  tuyeres  bright. 

Very  hot.    All  tuyeres  bright. 

Very  hot.     All  tu.veres  bright. 

Slag  hot  and  smoking     Tu_yeres  bright. 

Slag  hot  and  smoking.     Tuyeres  bright. 

Slag  hot  and  smoking.    Tuyeres  bright. 

Slag  still  hot,  but  not  quite  so  strikingly 
so  as  with  lower  pressure.  Tuyeres  sat 
isfactory,  but  beginning  to  form  noses. 


Less  hot.    Decided  noses. 



Much  cooler.  All  tuyeres  require  opening. 

*  Normal  pressure  and  slag  assay. 

These  tests,  although  not  entirely  uniform  in  every  respect,  are 
still  quite  regular  and  agree  closely  with  many  previous  observations. 

With  the  highest  available  blast,  14  ounces  per  square  inch,  the 
production  still  increases,  though  only  slightly  above  the  normal 
capacity,  but  it  is  evident  more  wind  is  introduced  than  can  be 
consumed  by  the  fuel;  a  lowering  of  temperature  occurs,  a»  dis- 
tinctly shown  by  the  appearance  of  the  slag;  and  thick,  hard  noses 
are  formed  about  each  wind  stream,  which  would  soon  obstruct 
the  blast,  and  probably  cause  a  general  chilling  of  the  furnace. 

Judging  from  this  series  of  tests,  as  well  as  from  numerous 
former  trials,  when  smelting  both  lead  and  copper  ores  of  many 
different  varieties  in  cupolas  of  various  sizes  and  under  very  vary- 
ing conditions,  it  seems  advisable  to  limit  the  blast  pressure  to  the 
point  just  indicated  except  where  furnaces  are  to  be  used  simply 
for  melting,  regardless  of  any  possible  oxidizing  effect.  In  no 
single  instance  has  anything  more  than  a  temporary  increase  of 
capacity  accompanied  a  blast  pressure  above  12  ounces  per  square 
inch,  and  the  rapid  cooling  of  the  furnace  and  formation  of  heavy 
and  solid  noses  have  soon  brought  the  experiment  to  a  termination. 

It  seems,  therefore,  that  a  pressure  of  from  8  to  12  ounces,  with 
a  tuyere  diameter  of  from  3  to  5^  inches,  is  best  suited  to  the  ordi- 
nary conditions  of  copper  smelting.* 

The  employment  of  soft-wood  charcoal  or  other  fragile  fuel  may 
make  it  necessary  to  diminish  even  this  light  pressure,  while  an- 
thracite may  demand  a  more  powerful  blast  for  its  most  econom- 
ical use. 

*  These  words  were  written  for  tbe  earlier  editions  of  this  book,  and  since  that 
time  experience  has  taught  thn  important  effect  produced  by  blast  pressure 
upon  the  oxidizing  intiuence  of  the  blast-furnace. 


I  have  mauy  times  used  wood  in  two-foot  ieugths  to  replaet;  a 
portion  of  the  coke  in  the  blast-fnruace,  though  merely  to  tide 
over  a  time  when  coiie  was  scarce.  I  have  invariably  noticed  a 
decided  rise  in  the  grade  of  the  matte  when  smelting  with  wood. 
Mr.  Herbert  Lang*  gives  some  interesting  information  on  the 
subject,  as  follows: 

"CordwGod,  sawn  in  blocks  of  a  foot  in  length,  is  a  regular  con- 
bjitueut  of  our  fuel  charge  at  Mineral,  Idaho,  our  work  being  the 
matting  of  silver  ores  by  fusion  in  a  blast-furnace.  The  furnace 
is  a  round  water-jacket  furnace,  of  30  inches  diameter  at  the 
tuyeres,  and  the  charge  of  smelting  mixture  weighs  950  pounds, 
requiring  110  pounds  of  Oonnelisville  coke  to  drive  it.  I  re])lace 
half  of  this  coke  with  135  pounds  of  firwood,  cut  from  dead  and 
apparently  perfectly  dry  trees.  This  mixture  produces  as  high  a 
smelting  temperature  as  all  coke,  whence  1  infer  that  the  smelting 
eifect  of  a  given  weight  of  wood  is  to  that  of  the  same  weight  of 
coke  as  11  to  27,  orl  to  2y\-.  A  cord  of  wood  sawed  ready  for  use 
weighs  2,340  pounds,  costs  15,  and  is  equivalent  to  8G6  pounds  of 
Oonnelisville  coke,  which,  at  125  per  ton,  costs  110.92,  or  rather 
more  than  twice  as  much  as  wood  per  unit  of  smelting  power.  The 
saving  by  the  use  of  wood  plus  coke,  over  coke  alone,  is  therefore 
To  cents  per  ton  of  ore.  The  principal  advantage,  however,  is  not 
in  the  saving  of  cost,  but  in  the  fact  that  a  great  deal  of  sulphur 
is  burned  off  by  the  wood,  thus  allowing  the  use  of  a  greater  pro- 
portion of  sulphide  ores  in  the  charge,  which  is  a  point  of  great 
moment,  as  such  ores  predominate  here,  and  we  are  as  yet  unpro- 
vided with  roasting  apparatus.  To  offset  these  advantages,  the 
wood  produces  a  great  deal  more  flue-dust — twice  as  much,  I 
should  think — and  reduces  the  smelting  capacity  about  one-third. 
With  the  fuel  mixture  described,  I  can  carry  only  six  ounces  of 
blast;  but  the  furnace  keeps  in  good  condition  above  and  below, 
the  tuyeres  remain  unatfected,  the  slag  is  hot  and  reasonably  free 
from  valuable  metals,  and  the  conditions  of  successful  smelling 
are  met  in  all  respects,  except  as  to  the  serious  reduction  of  tonnage. 

"Mr.  Dwight,  in  his  comments  upon  Mr.  Neill's  jjaper  on  'The 
Use  of  Stone  Coal  in  Lead  Smelting,''  appears  to  infer  that  the 
coal  has  to  be  converted  into  coke  inside  the  furnace  before  it  can 
perform  useful  work.  1  presume  he  would  also  infer  that  wood 
has  to  become  charcoal   before  it  can   do  its  smelting  work,  but 

*  Transactions  American  Institute  Mining  Engineer's,  Vol.  XX.,  p.  545. 


that  such  an  iufereuce  is  erroneous  appears  from  the  fact  that  our 
firwood  produces  but  about  '20  per  cent,  of  charcoal,  and  that  of  a 
very  poor,  fragile  sort.  Accordiugly,  135  pounds  of  wood  would 
produce  only  27  pouuds  of  charcoal,  a  quantity  clearly  insufficient 
to  replace  55  pouuds  of  coke.  I  therefore  believe  that  the  volatile 
constituents  of  the  wood  do  a  considerable  amount  of  useful  work 
in  the  smelting  before  escaping  from  the  furnace.  The  smoke, 
which  is  very  thick  and  abundant,  has  a  peculiar  nauseating  odor, 
giving  no  evidence  of  free  snlpliurous  acid — a  circumstance  whicii 
leads  me  to  believe  that  the  sulphur  so  largely  burned  ot!  forms  a 
volatile  compound  with  the  organic  matters  sublimed  from  the 
wood,  the  reaction  perhaps  iurnishiug  a  considerable  amount  of 
heat.  I  presume  tliat  the  use  of  denser  kinds  of  wood,  such  as 
mountain  mahogany,  oak,  hickory,  ash,  etc.,  would  give  still  better 

Mr.  James  W.  Xeill,  of  Leadville,  Colorado,  has  made  some  ex- 
periments on  the  use  of  bituminous  non-coking,  and  semi-coking 
coal  in  the  lead-silver  blast-furnace,  that  are  highly  suggestive  to 
copper  smelters.     I  quote  from  his  paper:* 

"Bituminous  coal  has  for  many  years  been  used  for  the  smelting 
of  iron  ores  in  the  blast-furnace.  In  some  districts  in  Scotland  it 
is  used  alone,  in  others  it  is  used  mixed  with  coke.  The  similar 
use  of  certain  bituminous  coals  in  the  United  States  has  been  re- 
peatedly mantioned.  In  the  lead-silver  smelting  blast-furnace, 
however,  tiie  requirements  of  iron  smelting  are  not  present.  Here 
the  general  question  is,  which  fuel,  or  fuel  mixture,  will  permit 
the  most  rapid  driving  of  the  furnace?  Tbe  conditions  of  efficiency 
in  the  reduction  of  iron,  of  sufficient  heat,  of  capacitv  to  carry  the 
burden,  etc.,  are  usually  satisfied  by  anv  of  the  commercial  cokes 
of  the  regions  surrounding  the  lead-silver  smelting  districts,  and 
it  is  therefore  usually  the  price  which  decides  the  choice  of  coke. 
In  most  of  these  smelting  districts,  bituminous  coal  of  non-coking 
character  is  very  much  cheaper  tlian  the  coke,  and  its  use  alone,  or 
with  coke,  would  materially  lessen  the  fuel  expense  per  ton  of  ore. 
This  saving,  if  achieved  without  occasioning  other  losses  in  the 
working  of  the  furnaces,  would  be  net  gain  to  the  silver-lead 
smelter,  and  the  following  experience  with  the  use  of  bituminous 
fuel  is  given  to  show  what  can  be  done  in  this  way. 

"About  1884,  while  in  charge  of  the  smelting  works  at  Mine 

*  7 raasactions  Aineriean  Institute  Afining  Engineers,  Vol.  XX.,  p.  165. 


La  Motte,  Missouri,  1  ran  short  of  coke,  aud  haviug  a  supply  of 
stone  coal  on  hand,  I  replaced  half  of  the  coke  in  the  charge  with 
this  coal,  continuing  its  use  until  a  supply  of  coke  arrived.  Dur- 
ing this  period  (about  twenty-four  hours)  no  noticeable  change  in 
the  working  of  the  furnace  occurred;  but  as  the  stone  coal  was 
more  expensive  than  the  coke,  the  practice  was  not  continued. 
With  the  precedent  of  various  authorities,*  aud  my  own  brief  per- 
sonal experience,  and  in  view  of  the  circumstance  that  in  Leadville 
to-day  coke  costs  three  times  as  much  as  certain  kinds  of  coal,  I 
have  recently  ventured  to  experiment  with  Rocky  Mountain  coal 
in  the  blast-furnaces  of  the  Harrison  Keductiou  Works. 

"These  furnaces  are  78  by  36  inches  in  size  at  the  tuyere  line, 
have  10  inches  bosh  in  the  jackets,  and  are  about  12  feet  high 
from  tuyere  to  charge  door.  At  the  time  the  experiments  com- 
menced, the  furnaces,  haviug  been  running  some  time,  were  in 
bad  condition  from  zinc  accretions  in  the  upper  part  of  the  stacks, 
aud  would  have  to  be  "blown  down"  and  "barred  out"  in  a  few 
days  at  furthest;  and  I  reflected  that,  if  the  coal  should  prove 
impracticable  as  fuel,  this  event  would  only  be  slightly  hastened. 

"On  January  20,  1891,  I  replaced  on  No.  1  furnace  50  pounds 
of  Cardilf  coke  with  60  jjounds  of  lump  coal  of  a  non-coking  vari- 
ety. The  fuel  charge  before  the  change  had  been:  Coke,  185 
pounds;  charcoal,  65  pounds.  It  was  now:  Coke,  135  pounds; 
stone  coal,  60  pounds,  and  charcoal  65  pounds.  The  charge  was 
made  on  the  evening  of  the  20th.  Next  morning  showed  no  ap- 
preciable change;  slag  assays  were  good,  but  in  the  afternoon  the 
slag  commenced  to  thicken  and  get  colder,  and  finally  refused  to 
run  out  of  the  tap-hole,  filled  all  the  tuyeres,  and  compelled  a 
stoppage  of  the  furnace.  On  taking  out  the  tap-jacket,  I  found 
that  the  charge  had  slipped  down,  filling  the  basin  with  raw  mate- 
rial, which  had  stopped  the  slag.  We  heaved  out  a  quantity  of 
this  raw  material  and  cleaned  the  tuyeres,  and  put  on  the  blast 
again,  when  the  furnace  cleared  itself  without  further  serious 
trouble.  Much  to  our  surprise,  we  found  the  tojj  of  the  furnace 
in  better  condition,  the  amount  of  accretions  hanging  on  the  sides 
being  much  less  and  the  charges  settling  more  evenly. 

"On  the  22d,  as  No.  1  continued  to  do  well  on  this  fuel,  I  put 

*  T.  Sterry  Hunt.  Transnctions  Ammcan  Ini^titute  Mining  Engineers,  Vol. 
II.,  p.  275;  Vol.  VII..  p.  313:  J.  S.  AlpxanHer.  Vol.  I.,  p.  225;  A.  Eilser,  Vol. 
I.,  p.  216;  Phillip.  Elements  of  Metallurgy,  p.  250;  Groves  &  Thorpe,  Chemical 
Technology,  Vol.  I. 


the  same  fuel  charge  upon  furnace  No.  4,  the  upper  part  of  which 
was  also  in  bad  condition.  Here  the  result  was  the  same;  but  by 
careful  watching  serious  trouble  was  avoided.  Meanwhile^  instead 
of  our  having  to  clean  No.  1  in  a  day  or  two,  it  ran  13  days  after 
the  coal  was  first  put  on. 

''  During  the  first  week  in  February,  all  the  furnaces  were  blown 
down  and  barred  out,  and  blown  in  again  on  the  above  fuel  charge, 
all  starting  off  nicely.  On  the  4th  of  February  the  charge  was 
changed  to:  Coke,  120  pounds;  stone  coal,  70  pounds;  charcoal, 
TO  pounds;  and  on  this  charge  they  ran  until  the  14th.  During 
this  period  the  coal  used  was  a  lump  coal  from  Bouse,  Colorado,  a 
semi  coking  coal  of  good  quality,  a  sample  giving  us  8.91  per  cent, 
of  ash.  The  pecuniary  saving  was,  that  in  14  days  we  replaced 
126.98  tons  of  coke  at  $8,  costing  $1,015.84,  with  148.15  tons 
lump  coal  at  15.50,  ousting  1814.82,  a  difference  of  $201. t)2. 

"On  February  14th  I  replaced  the  lump  coal  with  an  equal 
weight  of  pea  coal  from  the  Sunshine  mines.  A  sample  of  this 
gave  us  7.72  per  cent,  of  ash;  it  is  a  semi-coking  coal,  and  costs 
us  $2.50  per  ton.  The  furnaces  ran  for  the  remaining  14  days  on 
the  same  charge  of  fuel  as  during  the  first  half  of  the  mouth, 
doing  their  work  nicely.  During  this  time  we  replaced  126.84 
tons  of  coke  at  $8,  or  $1,014.72,  with  147.98  tons  of  pea  coal  at 
$2.50,  or  $369.95,  thus  saving  $644.77.  The  total  saving  for  the 
month  of  February  was,  therefore,  $845.79,  which  amounts  to 
about  26  cents  per  ton  of  material  smelted. 

"During  March  we  used  about  the  same  fuel  charge,  the  coal 
being  part  nut,  part  pea,  and  part  lump,  each  used  separately; 
and  the  total  saving  by  the  use  of  these  fuels  in  place  of  coke  was, 
for  March,  $1,075.08,  or  about  30  cents  per  ton  of  material.  For 
a  few  days  I  increased  the  amount  of  stone  coal  to  100  pounds, 
using  witli  it  100  pounds  of  coke  and  70  pounds  of  charcoal;  but 
the  furnace  did  not  work  well  on  this  charge,  the  hearth  becoming 
clogged  and  slag  flooding  the  tuyeres.  Whether  this  was  due  to 
the  fuel  or  to  bad  charging  remains  to  be  proved  by  further  tests. 

"Aside  from  the  direct  saving  by  the  use  of  the  stone  coal,  we 
have  observed  the  following  advantages  in  the  working  of  the 

^^  Below. — The  slags  appear  better  reduced  and  hotter,  and  the 
matte  separates  very  well.  Slag  assays  have  been,  if  anything,  lower 
than  on  the  old  fuel;  and  this  is  particularly  the  case  when  the 
charges  are  hanging  above.      The   jackets  keep  hotter  and   the 


tuyeres  brighter;  thus  the  fuel  is  more  completely  cousunied;  the 
furnace  'rods'  more  easily;  the  crucible  keeps  opeu  better;  the 
lead  is  hotter. 

^'  Above. — The  volume  of  smoke  is  somewhat  increased  and  smells 
decidedly  'tarry,'  but  does  not  look  different  from  that  of  the  old 
fuel;  the  charges  settle  much  more  evenly  and  the  tire  does  not 
creep  to  the  top;  thus  the  furnaces  seldom  flame.  As  the  tops 
keep  much  cooler,  the  production  of  flue-dust  is  much  smaller, 
and  the  losses  of  metals  by  volatilization  must  also  be  diminished. 
The  furnaces  now  run  about  a  week  longer  than  formerly  before 
needing  to  be  barred  out;  and  this  operation  is  no  more  dithcult 
than  before;  but  on  blowing  them  down,  the  charge  now  sinks 
much  further  before  flaming  commences. 

"The  pressure  of  the  blast  has  not  changed  materially  from  tiiat 
of  the  old-fuel  charge,  but  now  remains  more  uniform.  The  nec- 
essary reduction  of  iron  can  thus  always  be  relied  on. 

"To  the  practical  furnace-man,  these  advantages