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Cornell  University  Library 

TN  780.G79 

The  hydrometallurgy  of  copper, 

3  1924  004  678  755 

Cornell  University 


The  original  of  tiiis  book  is  in 
tine  Cornell  University  Library. 

There  are  no  known  copyright  restrictions  in 
the  United  States  on  the  use  of  the  text. 


Published  by  the 

McGraw-Hill   Book^ Company 

Nevvf  Yoirlk. 

5ucce5sor«  to  theBookDepartments  or  the 

McGraw  F'ublishing  Company  Hill  Publishing  Company 

Putlishers  of  Dook^  for 
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Metallurgical  and  Chemical  Engineering  Rjwer 






PART  I.    R()ASTIX(; 





Copyright,  1912,  by  the 
McGhaw-Hill  Book  Company 



The  information  available  on  the  hydrometallurgy  of  copper  is 
somewhat  fragmentary  and  widely  scattered.  The  wet  methods,  for 
treating  copper  ores,  are  diverse;  as  yet,  the  industry  has  not  arrived  at 
any  established  practice,  and  it  is  questionable,  on  account  of  the  widely 
different  character  of  the  ores,  if  at  any  time  one  routine  practice  will 
succeed  in  eliminating  other  processes  entirely.  In  the  discussion  of  the 
various  methods  it  is  intended  to  cover  all  the  most  essential  phases  of  the 

Roasting,  both  oxidizing  and  chloridizing,  has  been  given  a  prominent 
place  in  the  book,  because  on  many  ores,  especially  the  sulphides,  hy- 
drometallurgical  processes  are  directly  or  indirectly  dependent  upon  this 
step  for  successful  treatment. 

The  hydrometallurgy  of  copper  differs  from  the  hydrometallurgy  of 
gold  and  silver  largely  on  account  of  the  greater  percentage  of  material 
recovered.  For  this  reason  the  discussion  of  the  precipitation  plays  an 
important  part.  The  commercial  success  of  any  particular  process  will 
frequently  depend  on  the  nature  of  the  precipitant  and  the  cost  of 

The  book  is  the  result  of  notes,  covering  a  long  period  of  time,  from 
various  sources  and  from  my  own  experimental  work.  It  is  intended,  in 
the  text,  to  give  full  credit  for  the  various  sources  of  information. 

William  E.  Greenawalt. 
Denver,  Colorado, 
August,  1912. 


Pkeface     V 



Prepaiiation  of  the  Ore 1 

Relation  of  copper,  gold  and  silver.  Preparation  of  the  ore — Dry  crushing 
with  rolls — Dry  crushing  with  ball  rolls. 


Fuel 6 

Wood— Oil— Coal. 


Oxidizing  Roasting 12 

Objects  of  roasting — Chemical  combinations  of  the  metals  before  roasting — 
General  chemical  reactions  during  roasting — Essential  factors  in  roasting — 
Time — Temperature — Valentine's  temperature  experiments — Air — Rab- 
bling— Effect  of  metallic  sulphides  if  heated  with  exclusion  of  air — Sulphur 
— Decomposition  temperature  of  the  various  sulphates — Amount  of  sulphur 
trioxide  (SO3)  in  the  sulphur  dioxide  (SO2)  escaping  from  roasting  furnaces — 
Sulphur  determinations — Tellurium — Copper — Silver — -Gold — Lead — Zinc — 
Arsenic  —  Roasting  argentiferous  cobalt-nickel  arsenides  —  Antimony — 
Bismuth — Nickel  —  Calcium  (lime) — Magnesium — Manganese — Aluminum 
— Barium.    Alkali  metals — Chlorine.    Bromine — Loss  of  weight  in  roasting. 


Chloeidizing  Roasting 63 

Object  of  chloridizing  roasting — Adaptability  of  the  various  ores  to  chloridiz- 
ing  roasting — Chemistry  of  chloridizing  roasting — Arsenic  and  antimony — 
Zinc — -Lead — Calcium  carbonate — Magnesium — Quartz — Barium  sulphate — 
Sodium  sulphate — Percentage  of  salt — Time  of  adding  salt — Heap  chloridi- 
zation — Composition  of  the  roasted  ore — Volatilization  of  the  silver — Vola- 
tilization of  the  gold — Chloridization  of  copper  ores — Principal  factors  in  the 
loss  of  silver  and  gold  by  volatilization — Temperature — Time — Air  or  oxy- 
gen— Experiments  as  compared  with  practice — Relation  of  sulphur  to  the 
chloridization  of  silver  and  gold — Determination  of  loss  by  volatilization — 
Chloridization  determination. 


Pyhombtry 79 

Color  names  of  temperatures — Pyrometric  determinations. 





Roasting  Furnaces 83 

Hand  reverberatories  —  Short  reverberatories  —  Long  reverberatories  — 
Method  of  operating  a  long  hand  reverberatory — Cost  of  roasting  in  long 
reverberatories — Modified  long  reverberatories — Mechanical  reverberatories 
— Cost  of  mechanical  reverberatories — Fuel  required  in  roasting — Hearth 
area  required  in  roasting  various  ores — The  Brown  furnace — The  Pearce 
furnace — The  Holthoff-Wethey  furnace — The  Merton  furnace — The  Edwards 
furnace — The  McDougal  furnace — The  Herreshoff  furnace — The  Wedge  fur- 
nace— The  Greenawalt  porous  hearth — Bruckner  furnace — Howell-White 
furnace — Muffle  furnaces — Ore  coolers — Dust. 


Typical  Examples  of  Roasting 146 

Roasting  of  Cripple  Creek  ores — Roasting  arsenical  sulphide  ore  at  the 
Golden  Gate  mill,  Mercer,  Utah — Roasting  of  Casilas  concentrates,  Victoria, 
Australia — Roasting  at  Kalgoorlie. 



Properties  and  Solubilities  of  Copper      155 

Copper — Influence  of  impurities  on  the  properties  of  copper — Cupric  car- 
bonate— Cupric  nitrate — Cupric  oxide — Cuprous  oxide — Cupric  sulphate — 
Cupric  chloride — Cuprous  chloride — Cupric  silicate — Cuprous  sulphide — 
Cupric  sulphide — Cupric  hydroxide — Copper  cyanides — SolubiUty  of  sulphur 


Hydrometallurgical  Processes 169 

Classification  and  general  consideration — Chemical  processes. 


Chemical  Alkali  Processes 172 

The  Mosher-Ludlow  ammonia-cyanide  process — Sulphite  processes — Neill 
process — Van  Arsdale  process — Sulphate  processes — ^Acid  plants  at  the 
mine — Sulphuric  acid  leaching  of  oxidized  copper  ores  at  Clifton,  Arizona — 
Leaching  plant  at  the  Snowstorm  mine — Copper  leaching  plant  at  the  Gu- 
meshevesky  mine,  Russia — ^Ferric  sulphate — Experiments  with  ferric  sul- 
phate at  Cananea — Thomas'  experiments  with  ferric  sulphate  on  sulphide  ore 
— Experiments  in  Southern  Tyrol,  Spain — Copper  extraction  at  Kedabeg, 
Russia — The  Millberg  process — The  Ehiott  process — The  Laist  process — 
Method  of  extracting  copper  at  Rio  Tinto,  Spain — Treatment  by  heap 
roasting  and  leaching — Elimination  of  arsenic,  antimony  and  bismuth — • 
Chloride  processes — Hydrochloric  acid— Ferric  chloride — Doetsch  process 
— The  Froelich  process — ^Ferrous  chloride  process  — The  ferrous  chloride 
process  as  carried  out  at  Ore  Knob,  Ashe  Co.,  N.  C. — Hunt  and  Douglas 
process — The  Bradley  process — Longmaid-Henderson  process  for  treating 



pyritic  cinders — Longmaid-Henderson  process  at  the  Helsingbprg  copper 
works,  Sweden — Longmaid-Henderson  process  at  works  of  the  Pennsylvania 
Salt  Manufacturing  Co. — Cost  of  producing  copper  by  the  Longmaid- 
Henderson  process,  having  mechanical  roasters — Extraction  of  copper  from 


Copper  Precipitants 270 

Iron — Sponge  iron — Crucible  method  of  iron  ore  reduction — Precipitation 
with  a  coke-iron  couple — Precipitation  with  a  copper-iron  couple — Hydrogen 
sulphide — -Lime. 


Electrolytic  Processes 283 

General  consideration  of  electrolytic  methods — Definitions — Anode — 
Cathodes — Diaphragms — Current  density — General  laws  governing  the  elec- 
trodeposition  of  copper — ^Faraday's  law — Theoretical  data  for  copper  deposi- 
tion— Loss  of  energy  in  electrolytic  work:  Joule's  law — The  electrolyte — 
Effect  of  bismuth,  arsenic,  and  antimony  in  the  electrolyte — Iron  in  the 
electrolyte — Purification  of  the  electrolyte — Depolarizers — Rapid  depo- 
sition of  copper — Sulphuric  acid  process — Electrolytic  extraction  of  copper 
from  ore  at  Medzianka,  Poland — ^Plant  of  the  Intercolonial  Copper  Co., 
N.  S.,  Canada — Keith  process  at  Arlington,  N.  J. — The  Siemens-Halske 
process — M.  De  Kay  Thompson,  Jr.'s  experiments  on  the  Siemens-Halske 
process^Siemens-Halske  process  in  Spain — Experiments  at  the  Ray  mines, 
Arizona — Tosizza  process — Ramen  process — Treatment  of  ore  at  the  Braden 
Copper  Go's,  mine,  Chile — Extraction  of  copper  from  matte — Marchese 
process — Gunther  process — Electrolytic  chloride  processes — Electrolysis  of 
oupric  chloride — -Electrolysis  of  cuprous  chloride — The  Body  process — The 
Hoepfner  process — The  Douglas  process — ^The  Greenawalt  process — The 
Swinburne-Ashcroft  process — Baker-Burwell  process. 


Extraction  of  Precious  Metals  prom  Copper  Ores 363 

Method  proposed  by  Bertram  Hunt — Cyaniding  of  cupriferous  gold  ore  of 
the  Bagdad  Gold  Mining  Co. — Extraction  of  precious  metals  with  chloride 
solutions — Electrolytic  chlorine,  theoretical  data — ^Practical  data— Treat- 
ment of  ore  at  Mt.  Morgan  mine — Treatment  of  auriferous  copper  ores  at 
Falun,  Sweden. 


Treatment  op  Zinciferous  Copper  Ores 382 

The  Hoepfner  zinc  process. 


Treatment  of  Copper- Nickel  Ore  and  Matte 387 

The  Hoepfner  nickel  process — The  Browne  process — Sjostedt- James  process 
— Gunther-Franke  process — Hybinette  process — Cito  process. 




Precipitation  of  Copper  from  Mine  Waters 

Precipitation  of  copper  from  mine  waters  at  Butte — Precipitation  of  copper 
from  mine  waters  at  Copper  Queen  Consolidated  Mining  Company,  Bisbee, 
Arizona — Leaching  of  copper  ore  in  place. 


Refining  of  Copper  Precipitate 

Oxidizing  stage — Reducing  stage — Furnaces  used  for  copper  refining. 


Copper  Sulphate,  Bluestone ^^" 

The  Oker  process — The  Freiberg  process — Hofmann's  copper  sulphate 


Apparatus  and  Appliances 450 

Tanks — Conducting  solutions — Regulating  the  flow  of  solution — Elevating 
solutions — Montejus  tanks — Air  lifts — Stoneware  pumps — Rubber  pumps — ■ 
Pressure  tanks — Ore  agitation  and  filtration — Filter  presses  for  filtering  acid 
copper  solutions — Construction  of  floors. 


Power  Data 473 

Comparison  of  steam  and  water-power  plants — Cost  of  water-power  in  the 
Western  United  States — Comparison  of  steam  and  producer  plants — Pro- 
ducer gas  plant  using  western  lignite — Consumption  of  fuel  in  a  gas 
engine — Cost  of  power  with  steam-plants — Oil  engines. 


Economic  Considerations 483 

Cost  of  depositing  copper  electrolytically — Comparison  of  electrolytic 
methods  with  the  iron  precipitation  process,  using  sulphuric  acid  as  the 
solvent — Economic  relation  of  current  density  and  voltage — Chemical 
treatment  of  sulphide  ores — Comparison  of  wet  methods  with  smelting — 
Chemical  and  mechanical  difficulties — General  applicability  of  hydrometal- 
lurgical  processes. 




Relation  of  Copper,  Gold,  and  Silver.— Copper,  gold  and  silver  are 
chemically,  mineralogically,  and  metallurgically  intimately  associ- 
ated. Chemically,  they  occur  in  the  same  group  in  the  Periodic  System; 
mineralogically,  one  of  these  metals  is  rarely,  if  ever,  found  unaccom- 
panied by  one  or  both  of  the  others;  and  metallurgically,  any  scheme 
which  contemplates  the  profitable  recovery  of  the  copper  must  take  into 
consideration  the  profitable  recovery  of  the  accompanying  precious 
metals  also. 

In  the  hydrometallurgical  treatment  of  copper  ores,  or  of  gold  and 
silver  ores  containing  copper,  it  is  evident  that  the  extraction  of  all 
three  metals,  concurrently  or  consecutively,  must  be  given  due  consider- 
ation. Many  operations  in  the  wet  treatment  of  ores  of  one  of  these 
metals  are  applicable  to  the  others.  In  the  acid  treatment  of  gold  and 
silver  ores,  as  in  chlorination,  the  conditions  of  roasting  and  extraction 
are  not  essentially  different  from  the  treatment  of  copper  ores  with  acid 

Many  ores  may  be  treated  by  hydrometallurgical  processes  without 
roasting.  Others,  especially  the  sulphides,  have  to  be  roasted,  and  the 
sulphides  constitute  the  greater  proportion  of  the  available  ores  of  copper. 
Some  sulphides  may  be  treated  without  roasting,  but  such  treatment  is 
the  rare  exception,  and  for  obvious  reasons  will  probably  not  find  exten- 
sive application. 

The  best  conditions  of  roasting,  for  the  various  ores  for  chemical 
treatment  by  any  of  the  solvent  processes,  are  very  much  the  same, 
whether  the  metals  to  be  extracted  are  copper,  gold  or  silver,  or  all 
combined  in  the  same  ore.  The  same  furnaces  are  used;  the  same  costs 
apply;  and  the  conditions  of  roasting  which  give  the  best  extraction  of  the 
precious  metals  will,  in  general,  also  give  the  best  extraction  of  the  copper. 

In  roasting  copper,  gold  and  silver  ores,  for  hydrometallurgical  treat- 
ment the  metals  themselves  offer  no  particular  difficulty  in  the  operation. 
The  difficulties  encountered  in  roasting  will  usually  be  in  the  nature  of 



the  other  elements  associated  with  them  in  the  gangue.  It  is  evident 
that  in  considering  the  roasting  of  copper  ores,  or  copper  ores  containing 
gold  and  silver,  the  foreign  elements  must  be  taken  into  account  quite 
as  seriously  as  the  metals  themselves. 

Roasting  of  ores,  as  a  step  for  their  treatment  by  solvent  processes, 
is  materially  different  from  that  required  for  subsequent  smelting. 
While  the  chemical  reactions  during  the  roasting  are  essentially  the  same 
for  both  methods,  a  good  roast  for  a  solvent  process  requires  vastly 
more  delicate  manipulation  and  a  more  thorough  elimination  of  the 
sulphur.  A  roast  which  would  be  satisfactory  for  smelting  might  be, 
and  usually  is,  absolutely  worthless  for  treatment  by  wet  methods;  on 
the  contrary,  ore  which  is  satisfactorily  roasted  for  treatment  by  wet 
processes  would  be  satisfactory  for  smelting  also,  but  the  expense  of 
roasting  would  be  considerably  greater. 

In  ores  containing  copper,  gold  and  silver,  if  the  precious  metals 
are  not  extracted  simultaneously  with  the  copper,  the  roasting  of  the  ore 
to  make  their  subsequent  extraction  satisfactory,  either  by  cyanidation 
or  chlorination,  must  be  taken  into  account. 

Cupriferous  pyritic  ores,  high  in  sulphur,  are  sometimes  roasted  in 
heaps,  preparatory  to  extracting  the  copper  as  soluble  sulphate,  but 
this  practice  is  not  finding  extended  application,  and  at  the  mine  where 
it  was  largely  employed  its  use  has  been  discontinued. 

The  only  roasting  which  is  finding  favor  for  the  hydrometallurgical 
processes  is  in  suitable  furnaces,  usually  reverberatories,  and  preparatory 
to  which  the  ore  is  crushed  fine  enough  to  be  thoroughly  roasted  in  sev- 
eral hours. 

Preparation  of  the  Ore. — Ore,  as  roasted  in  furnaces  for  hydromet- 
allurgical treatment,  is  usually  crushed  to  a  fineness  varying  from  8  to 
40  mesh.  Below  8  mesh  the  particles  become  too  large  for  efficient 
oxidation,  and  above  40  mesh  the  dust  is  likely  to  give  trouble.  On 
the  whole,  ore  ground  to  a  fineness  varying  from  10  to  30  mesh  will  give 
the  best  average  results. 

When  roasting  constitutes  a  step  in  any  metallurgical  process,  the 
ore  is  crushed  dry.  Rolls  and  ball  mills  are  best  suited  for  this  work. 
Concentrates  are  usually  the  product  of  wet  crushing.  If  the  ore  is  of 
too  low  a  grade  to  admit  of  direct  chemical  treatment,  concentration 
offers  a  means  of  increasing  the  tenor  of  the  material,  while  at  the  same 
time  eliminating  the  most  injurious  elements.  In  this  way,  lime  par- 
ticularly may  be  largely  eliminated  from  the  iiltimate  product  to  be 
treated  by  roasting  and  a  chemical  process.  If  concentration  forms  a 
step  in  the  general  treatment,  there  is  no  need  of  close  work  to  obtain 
a  high  grade  concentrate,  and  hence  there  is  no  need  of  excessive  loss  in 
the  tailings.  A  concentrate  containing  10  per  cent,  copper  would  be  a 
very  desirable  product  for  roasting  and  for  treatment  by  a  solvent  proc- 


ess,  and  such  a  concentrate  should  be  made  without  excessive  loss; 
whereas  if  shipment  to  a  smelter  is  desired,  such  a  product  would  not  pre- 
sent much  advantage,  and  to  get  a  higher  grade  product  would  result 
also  in  getting  a  considerably  greater  loss. 

The  moisture  contained  in  the  ore  before  charging  into  the  furnace 
should  also  be  considered.  Concentrates  may  be  charged  in  a  hand- 
rabbled  furnace  without  drying,  but  in  mechanical  roasters  it  is  evi- 
dently better  to  remove  the  moisture  sufficiently  so  that  the  ore  may  be 
fed  uniformily  into  the  furnace  by  mechanical  means.  Much  moisture 
in  the  ore  charged  into  the  furnace  has  a  tendency  to  cool  it  unduly. 

The  moisture  in  the  ores,  as  well  as  the  moisture  in  the  fuel  gases, 
has  an  important  bearing  on  the  chemistry  of  the  roasting  process. 

The  principal  expense  in  the  preparation  of  the  ore  for  roasting  is  in 
crushing.  This  may  vary  within  wide  limits,  depending  on  the  character 
of  the  ore,  the  fineness  to  which  it  is  reduced,  and  the  amount  crushed. 
Usually  it  will  vary  from  25  to  50  cents  per  ton,  with  a  reasonably  large 

Dry  Crushing  with  Rolls. — Rolls  are  largely  used  to  crush  ore  to 
medium  fineness.  For  grinding  finer  than  20  mesh  they  are  inferior  to 
some  other  type  of  machines,  and  it  is  a  question  whether,  under  any 
conditions,  they  are  as  satisfactory  as  ball  mills. 

It  costs  from  25  to  30  cents  per  ton  to  crush  Cripple  Creek  ore  to 
12  or  14  mesh,  on  a  basis  of  from  200  to  300  tons  per  day.  In  one  plant, 
having  two  48-in.  roughing  rolls  and  four  48-in.  finishing  rolls,  300  tons 
of  ore  are  regularly  crushed  per  day.  The  roughing  rolls  are  run  only 
during  the  daytime,  but  the  finishing  rolls  are  run  continuously  for 
three  eight-hour  shifts,  with  three  men  on  a  shift.  One  man  attends 
to  the  screens. 

In  another  plant,  having  one  36-in.  roll,  and  three  26-in.  rolls,  175 
tons  are  crushed  in  24  hours  to  from  30  to  40  mesh. 

A  combination  of  one  36-in.  roughing  roll;  one  26-in.  roll  doing 
medium  work,  and  two  26-in.  rolls  doing  finishing  work,  will  crush  from 
125  to  175  tons  of  ore  of  ordinary  hardness  to  30  mesh;  200  to  250  tons 
to  20  mesh,  and  250  tons  and  more  to  16  mesh. 

For  the  best  working  of  a  roll  crushing  plant,  it  is  essential  that  the 
reduction  shall  be  gradual  in  going  from  one  roll  to  the  next. 

Dry  Crushing  with  Ball  Mills. — ^For  dry  crushing  ball  mills  present 
certain  advantages  over  rolls  in  that  they  are  self-contained  and  the 
screening  is  simplified.  Their  capacity  is  also  large.  A  No.  5  Krupp 
ball  mill  will  crush  43  tons  daily  of  ordinary  sulphide  ore,  and  the  No.  8, 
100  tons;  using  from  18  to  23  h.  p.  for  the  5-ton  mill,  and  from  60  to  65 
for  the  8-ton  mill,  or  2.1  and  1.6  tons  per  horse-power  respectively. 

M.  W.  Von  Bernewitz'  has  given  some  valuable  information  on  ball 

'  Min.  and  Scientific  Press,  July  15,  1911. 


mill  practice  at  Kalgoorlie.     The  following  table  gives  a  summary  of  the 
essential  facts. 


of  plant 


Associated  Northern. 


Great  Boulder 


Perse  verence 

South  Kalgurli 

No.  of 

4  No.  8 
6  No.  5 
3  No.  5 

3  No.  5 

4  No.  8 
9  No.  5 
8  No.  8 
3  No.  8 
1  No.  5 

of  balls 




r.  p.  m. 


h.  p. 




Steel  con-  [   Life  ot 
sump-        grinding 
tion  plates 

lb.  per  ton 








Other  items 



of  plant 








Associated  Northern   . . 


Chaffers  .... 



0 .  435 




South  Kalgurli 



One  man  per  shift  of  8  hours,  at  lis.  8d.  (S3. 22)  can  look  after  eight 
No.  8  ball  mills;  but  in  smaller  plants  the  mill  man  attends  to  conveyors, 
elevators,  dust,  pipes,  etc.  Ball  mills  should  be  fed  with  no  larger  ore 
than  will  pass  a  3-in.  ring.  With  fine  ore  the  balls  are  likely  to  bed. 
Actual  weighing  has  shown  a  3-in.  feed  is  crushed  faster  and  the  wear  of 
steel  is  less  than  when  1-in.  material  is  fed  with  a  quantity  of  fines. 

A  No.  5  mill,  including  foundations,  may  be  erected  for  £600  ($2922.00) 
and  a  No.  8  for  £1000  ($4870.00),  bin  and  conveyor  inoluded.  It 
will  usually  cost  about   £300  ($1461.00)  per  year  for  upkeep  of  a  No.  5. 

At  the  Golden  Cycle  mill,  at  Colorado  Springs'  four  No.  66  "Kominu- 
ters"  have  a  capacity  of  17,000  lb.  of  Cripple  Creek  ore  per  hour  for 
each  mill,  when  fed  with  a  product  that  had  been  reduced  by  rolls  to 
pass  through  a  revolving  screen  made  of  1/4  in.  steel  plate  and  having 
openings    11/2  in.  in  diameter.     The  kominuters  were  equipped  with 

'  Loohiel  M.  King,  Mining  and  Scientific  Press,  Jan.  25,  1908. 


a  diagonal  slotted  sciceii,  size  of  opening  5/32  by  1/2  in.  No.  8  steel 
plate.  This  opening  gave  a  product  varying  in  size  from  1/8-in.  cubes 
to  the  finest  slimes-.  The  consumption  of  power  was  50  h.  p.  at  a  speed 
of  22  r.  p.  m.,  the  ball  consumption  being  fourteen  5-in.  forged  steel  balls 
weighing  about  19.5  lb.  each,  per  day  of  24  hours.  One  man  can  attend 
to  six  mills. 

The  average  results  from  several  types  of  ball  mills  show  that  one 
ton  of  steel  balls  will  crush  about  50  tons  of  ore  during  24  hours  from  a 
feed  1  1/2  in.  diameter  down  to  a  product  of  from  12  to  20  mesh  in  one 



Roasting,  as  a  step  in  the  treatment  of  ores  by  the  hydrometallurgical 
processes,  is  usually  carried  out  in  the  immediate  vicinity  of  the  mmes. 
At  industrial  centers,  the  consideration  of  fuel  is  a  very  simple  matter, 
but  not  so  in  copper,  gold  and  silver  mining  districts  where  the  selection 
of  a  particular  kind  of  fuel  is  frequently  a  matter  of  necessity,  based  on 
local  conditions.  In  vicinities  where  wood  is  abundant  it  will  ordinarily 
be  used  in  preference  to  the  more  expensive  coal,  which  has  to  be  freighted 
in.  If  a  mining  district  has  no  wood,  and  is  some  distance  from  the 
source  of  fuel  supply,  the  greater  calorific  power  of  oil  per  unit  of  weight 
over  that  of  the  coal  might  make  it  the  cheaper  fuel  on  account  of  the 
difference  in  cost  of  freight. 

Other  things  being  equal,  the  relative  desirability  of  fuels  for  roast- 
ing purposes  is  gas,  oil,  bituminous  coal,  wood,  lignite.  Anthracite  is 
not  often  available,  but  if  it  is,  there  is  a  decided  advantage  in  first 
converging  it  into  producer  gas. 

Most  of  the  expense  of  roasting,  in  mechanical  furnaces,  is  in  the 
fuel.  Some  of  the  essential  facts  pertaining  to  the  various  fuels  and  a 
comparison  of  their  relative  value  is,  therefore,  pertinent. 

Wood. — It  frequently  happens  that  in  mining  camps  far  removed 
from  coal  supply,  wood  can  be  obtained  cheaply  and  in  large  quantities. 
For  roasting,  if  the  wood  is  perfectly  dry,  it  is  more  desirable  than 
lignite  of  the  inferior  qualities  of  bituminous.  Green  wood  contains 
from  30  per  cent,  to  40  per  cent,  moisture.  After  thorough  seasoning, 
for  about  a  year,  in  the  open  air,  the  moisture  is  from  20  to  25  per  cent. 
The  wood  of  various  trees  are  nearly  identical  in  chemical  composi- 
tion, which  for  perfectly  dry  wood  and  of  ordinary  fire  wood  holding 
hygroscopic  moisture,  is  practically  as  shown  in  the  table  on  the  follow- 
ing page. 

The  ash  in  most  woods  varies  from  0.5  to  1.5  per  cent.  Most  of  the 
pines  and  others  of  the  coniferous  family  contain  hydrocarbons  (pitch, 
turpentine)  which  increase  their  value  as  fuel. 

In  steam-boiler  tests  wood  is  assumed  as  0.4  of  the  value  of  the 
same  weight  of  coal.  It  is  safe  to  assume  that  2  1/4  lb.  of  dry- wood  is 
equal  to  1  lb.  of  average  quality  bituminous  coal,  and  that  the  fuel 
value  of  the  same  weight  of  different  woods  is  nearly  the  same.  That 
is  to  say,  a  pound  of  pine  is  worth  as  much  for  fuel  as  a  pound  of  hickory, 
supposing  both  to  be  dry. 



Dessicated  wood 

Ordinary  fire  wood 


50  per  cent.                |             37 . 5    per  cent. 

6  per  cent.                              4.5    percent. 

41  per  cent.                            30.75  per  cent. 

1  npr  ppnt                                            H    7^  npr  ppnt. 





2  per  cent. 

1 . 5    per  cent. 

Hygroscopic  water 

100  per  cent. 

75 . 0    per  cent. 
25.0    percent. 

100 . 0    per  cent. 

It  is  important  that  the  wood  be  dry,  as  each  10  per  cent,  of  moisture 
in  wood  will  de1>ract  12  per  cent,  from  its  value  as  fuel. 

A  cord  of  wood  is  a  pile  4  ft.  by  4  ft.  by  8  ft.  which  is  equal  to  128  cu. 
ft.     About  56  per  cent,  is  solid  wood,  and  44  per  cent,  spaces. 

Fire-boxes  for  burning  wood  should  be  built  so  as  to  contain  a  deep 
bed  of  fuel.  They  should  be  narrower  at  the  bottom  than  at  the  top. 
With  properly  designed  fire-boxes,  burning  thoroughly  dry  wood,  a 
very  intense  heat  can  be  obtained  which  is  quite  as  effective  in  roasting 
ores  as  most  coals  available  in  copper  mining  districts. 

Where  wood  is  abundant  in  the  Rocky  Mountain  Region  it  will 
ordinarily  cost  from  $3.00  to  $3.50  per  cord,  cut  and  piled  at  the  metallur- 
gical works  ready  for  use. 

Wood  burns  with  a  long  flame  and  makes  comparatively  little  smoke, 
which  are  ideal  conditions  for  roasting. 

Charcoal  gives  out  much  more  useful  heat  than  wood,  because  the 
water  contained  in  the  wood,  or  formed  by  the  combustion  of  its  oxygen 
and  hydrogen,  has  to  be  evaporated  during  its  combustion.  100  parts 
of  wood  give  only  as  much  heat  as  40  parts  of  charcoal. 

Charcoal  is  made  by  the  dry  distillation  of  wood,  at  a  temperature 
of  from  460°  to  450°  C.  This  may  be  done  in  heaps  or  in  closed  retorts. 
Dry  wood  in  stacks  yields  about  one-fourth  its  weight  in  charcoal. 
Charcoal  develops  on  burning  8000  heat  units,  while  wood,  dried  in  the 
air,  does  not  develop  more  than  2800  units  of  heat.  Therefore,  seven 
parts  of  charcoal  gives  as  much  heat  as  20  parts  of  wood,  but  the  20  parts 
of  wood  are  capable  of  yielding  only  five  parts  of  charcoal. 

If  wood  has  to  be  transported  any  considerable  distance  for  roasting, 
it  might  be  profitable  to  convert  it  into  charcoal  at  the  forests  and  then 
burn  it  in  the  roasting  furnace,  after  having  converted  it  into  producer 

The  weight  of  a  bushel  of  charcoal  is  usually  taken  as  20  lb. 


One  cord  of  wood 
(128  cu.  ft.) 

Weight  in  pounds 
per  cord 

Pounds  of  coal  equiva- 
lent to  one  cord  of  wood 

Pounds  of  oil  equivalent 
to  one  cord  of  wood 





1,800  to  2,000 
1,540  to  1,715 

1,300  to  1,450 

940  to  1,450 
800  to     925 


White  Oak 



Red  Oak I 


Black  Oak 


Chestnut > 


Pine. .  . , 


It  might  be  said  that  the  approximate  heating  value  of  wood,  coal 
and  oil  is:  2  cords  of  average  pine  =  l  ton  of  average  bituminous  coal  = 
13/8  tons  of  lignite  =  3  1/2  to  4  barrels  of  crude  oil. 

As  to  the  absolute  consumption  of  fuel,  in  roasting,  much  depends 
on  the  nature  of  the  ore,  the  amount  of  sulphur  in  the  raw  ore,  and  the 
extent  to  which  the  sulphur  is  eliminated. 

Oil.— Oil,  next  to  gas,  is  the  most  desirable  fuel  for  roasting  purposes. 
It  is  largely  used  where  it  can  be  obtained  cheaply  and  the  supply  is 
constant.  It  was  for  many  years  the  principal  fuel  used  in  roasting 
Cripple  Creek  ores.  Recently,  owing  to  the  uncertainty  of  the  supply, 
producer  gas  has  largely  displaced"  the  crude  oil  and  residuum.  In 
California,  where  large  oil  fields  have  lately  been  developed,  it  is  dis- 
placing wood  and  coal  in  the  roasting  of  stamp  mill  concentrates  for 

Fuel  oil  has  the  following  advantages  over  coal  and  wood  in  roasting 

Reduction  of  weight  of  fuel  by  50  per  cent. 

Reduction  of  bulk  of  fuel  by  30  per  cent. 

Reduction  of  labor  by  50  per  cent. 

Prompt  kindling  of  fire. 

Cleanliness  and  freedom  from  ash. 

No  loss  of  heat  by  useless  radiation,  as  in  the  coal  fire-box  where 
the  heat  and  products  of  combustion  are  introduced  through  the  top  of 
the  arch. 

Convenience  in  directing  and  controlling  the  heat. 

It  is  possible  to  get  with  it  either  a  long  rolling  flame,  or  an  intensely 
hot  local  flame. 

Oil,  as  it  is  used  in  roasting  ores,  is  sprayed  with  a  steam  jet  directly 
into  the  furnace,  either  through  the  sides  or  through  the  arch.  The 
steam  is  usually  kept  at  a  pressure  of  from  60  to  90  lb.  and  the  oil  at 
from  30  to  50  lb. 

FUEL  9 

Most  of  the  oil  sold  for  fuel  purposes  ranges  from  14°  to  20°  Baunie. 
Oil  is  usually  bought  by  measure  and  not  by  weight.  The  lighter 
gravity  oils  contain  more  heat  units  per  pound  than  the  heavier  oils, 
but  there  are  more  pounds  of  fuel  in  a  gallon  of  heavier  oils  than  in  a 
gallon  of  lighter  oils.  The  gravity  of  the  oils,  therefore,  is  not  a  matter 
of  much  consequence. 

A  U.  S.  gallon  of  oil  weighs  from  6.5  to  7.2  lb.  and  42  gallons  are  taken 
as  a  barrel.  Residuum,  that  is,  the  residue  of  crude  oil  after  the  volatile 
substances  have  been  driven  off  by  heating,  is  largely  used  as  fuel  for 
roasting  purposes. 

In  some  of  the  mills  treating  Cripple  Creek  ore  both  coal  and  oil  are 
used  in  the  same  furnace.  In  some  of  the  roasting  furnaces  coal  is  used 
at  the  cooler,  or  feed  end,  while  in  others  the  reverse  is  the  case.  The 
relative  quantity  of  coal  and  oil  used  also  varies  greatly.  The  average 
consumption  might  be  considered  as  100  lb.  of  coal  and  15  gallons  of  oil 
per  ton  of  ore,  in  roasting  1  1/2  to  3  per  cent,  sulphur  down  to  about  0.5 
per  cent.  With  oil  alone,  it  takes  from  0.35  to  0.45  barrels  to  roast  a 
ton  of  ore,  in  addition  to  the  small  amount  of  fuel  necessary  to  generate 
the  steam  for  applying  the  oil. 

In  California  it  takes  about  half  a  barrel  of  oil  to  roast  a  ton  of  stamp 
mill  concentrates  suitable  for  chlorination,  and  about  50  lb.  of  coal 
to  furnish  the  steam  to  pump,  heat,  and  atomize  the  oil.  As  a  fuel  90 
gallons  of  California  oil  is  equal  to  1  ton  of  coal. 

If  fuel  has  to  be  transported  any  considerable  distance,  oil  offers 
advantages  in  the  cost  of  freight,  since  for  the  same  weight  it  has  about 
twice  the  heating  value  of  coal,  and  about  four  times  that  of  wood. 

Coal. — Coal  is  the  most  universally  used  fuel  in  roasting.  Its  quality, 
however,  varies  so  much  that  careful  investigation  of  the  different 
kinds  available  is  a  serious  matter.  A  long-flame  bituminous  coal,  if 
direct  firing  is  used,  is  the  best,  while  lignite,  with  its  short  flame  and 
low  heating  quality,  is  the  worst.  The  tendency  of  short-flame  coal  is 
to  give  an  intense  local  heat,  and  such  a  heat  is  highly  detrimental  to 
the  roasted  ore.  The  best  way  to  distribute  the  heat  is  either  to  gassify 
the  coal,  or  if  fired  direct,  get  what  is  known  as  a  semi-producer  action 
in  the  fire-box,  by  the  introduction  of  steam  and  air  under  the  grate.  By 
either  of  these  methods,  a  long  rolling  flame  may  be  obtained  in  the 
roasting  chamber. 

Any  coal,  wheter  anthracite,  bituminous,  or  lignite,  will  give  the 
most  satisfactory  result  by  being  first  converted  into  producer  gas,  and 
conducting  the  gas  from  the  producer  mains  into  the  different  parts  of 
the  furnace,  and  there  consuming  it,  so  that  the  atmosphere  shall  be 
highly  oxidizing  and  with  as  little  local  heat  as  possible.  This  is  best 
accomplished  by  introducing  the  gas  in  smaller  quantities  at  more  points 
in  the  roasting  chamber,  rather  than  in  larger  volumes  at  fewer  places. 



The  advantages  of  producer  gas  over  direct  firing  are: 

The  gas  may  be  produced  from  inferior  coal,  and  makes  more  avail- 
able heat  in  the  roasting  furnace  than  is  possible  with  any  coal  burned 
in  an  ordinary  grate. 

It  can  be  easily  introduced  into  the  roasting  furnace  at  any  pomt 
and  in  any  quantity  desired,  thereby  giving  a  diffused  heat  over  the 
entire  bed  of  the  ore. 

The  construction  of  the  arch  of  large  mechanical  furnaces  is  very 
much  simplified. 

The  producers  may  be  centralized,  so  that  the  handling  of  coal  and 
ashes,  by  mechanical  appliances,  may  be  greatly  facilitated. 

In  all  cases,  where  producer  gas  is  used  to  roast  ore,  the  air  necessary 
for  its  combustion  should  be  pre-heated.  This  can  be  done  at  the  least 
expense  by  an  air-heating  arrangement  in  the  furnace  dust  chamber. 

The  relative  average  value  of  the  several  classes  of  coal  may  be  ap- 
proximately determined  from  the  accompanying  tables. 

(Kent,  Min.  Ind.,  1900) 

Anthracite    and    semi-an- 


Bituminous — eastern 

Bituminous — western 


per  cent. 

per  cent. 



per  cent. 

1  to3 

8  to  12 

3  to  12 

1  to3 

3  to  10 

15  to  25 

1  to3 

3  to  15 

25  to  40 

4  to  14 

5  to  25 

35  to  50 

12  to  18 

5  to  25 

over  50 

per  cent. 

97  to  88 

75  to  85 

60  to  75 

50  to  65 

Less  than  50 

Heating  value 
b.  t.  u.  per  lb. 
of  combustible 

value  of  com- 
=  100 

14,700  to  14,900 

15,600  to  16,000 
14,800  to  15,200 
13,600  to  14,800 
11,000  to  13,000 



A  rough  estimate  of  the  relative  practical  value  of  the  several  classes 
of  coal  may  be  calculated  as  follows: 




b.  t.  u. 
per  lb. 







Relative  practical 

b.  t.  u. 

=  100 


















Bituminous — eastern 

Bituminous — western 


The  relation  of  the  heating  value  of  coal  to  its  ultimate  analysis 
may  be  estimated  by  Dulong's  formula,  usually  within  a  limit  of  error 
of  2  per  cent.     This  formula  with  average  figures  for  the  constants  is : 



Heating  value  per  pound  in  b.  t.  u.  equals: 

1  /  100[14,650C  +  62,000  (H  -  ^)  +  4000S] 

In  which  C,  H,  0  and  S  arre  respectively  the  precentages  of  carbon, 
hydrogen,  oxygen,  and  sulphur  in  the  coal. 

There  is  more  ash  in  the  smaller  size  coal  than  in  the  larger  sizes, 
due  principally  to  the  greater  quantities  of  dirt  and  slate,  as  shown  by  the 
following  analyses  of  different  sizes  of  anthracite. 

Size  of  coal 

Fixed  carbon 


Egg,  2.5  to  1.7  in 

Stove,  1 .  75  to  1 .  25  in 

Chestnut,  1 .  25  to  0 .  75  in . . . 

Pea,  0 .  75  to  0 .  50  in 

Buckwheat,  0 .  50  to  0 .  25  in . 

88 . 5  per  cent. 
83 . 7  per  cent. 
80.7  per  cent. 
79 . 0  per  cent. 
76 . 9  per  cent. 

5 . 7  per  cent. 
10.2  per  cent. 
12.7  per  cent. 
14 . 7  per  cent. 
16.6  per  cent. 


Objects  of  Roasting. — The  object  of  roasting  is  to  convert  the  ore 
into  a  condition  which  will  have  the  least  injurious  effect  on  the  chem- 
icals used,  and  to  simplify  their  application.  Roasting  is  essentiallj- 
oxidation.  Many  metallic  oxides  are  not  as  readily  attacked  by  the 
solvent  in  the  subsequent  chemical  treatment  as  the  metals  in  other 

The  solvents  for  copper,  gold  and  silver  are  among  the  most  ener- 
getic substances  known.  Chlorine,  for  example,  combines  with  those 
elements  with  which  oxygen  is  able  to  combine,  because  in  many  respects 
it  is  equally  if  not  more  energetic  than  oxygen,  and  replaces  it  in  the 
proportion  of  2  atoms  of  chlorine  to  one  of  oxygen:  Clji  0.  Chlorine 
cannot  displace  oxygen  from  many  of  its  oxide  combinations.  Iron 
is  universally  associated  with  copper,  gold  and  silver  ores  in  the  form 
of  oxide  or  sulphide.  Chlorine  very  rapidly  combines  with  iron  in  its 
sulphide  and  sulphate  combinations,  but  does  not  appreciably  displace 
the  oxygen  from  its  oxide  combinations.  Most  of  the  metals  are 
less  injurious  in  their  oxide  than  in  their  sulphide  combinations,  while 
others  are  not  much  , improved  by  the  change.  If  the  ore  is  to  be 
treated  for  its  copper  content  by  an  acid  process,  the  oxide  of  copper 
resulting  from  the  roasting  is  readily  soluble  in  either  hydrochloric 
or  sulphuric  acids,  while  the  sulphide  of  copper  is  quite  insoluble  in 
either  of  these  acids.  Calcium  is  acted  upon  by  chlorine  and  the 
acids  as  readily  in  its  oxide  as  in  its  carbonate  combinations.  Much 
of  the  calcium,  however,  in  roasting,  is  converted  into  the  sulphate, 
which  is  an  improvement,  since  it  is  practically  neutral  and  unaffected 
by  all  solvents. 

Many  injurious  metals,  such  as  arsenic,  antimony,  and  bismuth,  are 
volatile  at  a  high  temperature  and  are  expelled  during  the  roasting. 
Many  oxides  are  benefited  by  elevated  temperatures.  Dehydration 
agglomerates  the  particles  and  makes  a  better  leaching  product.  Ores 
containing  much  clay  and  talc  are  similarly  benefited.  And,  finally, 
roasting  makes  the  ore  particles  porous,  thereby  very  materially  increas- 
ing the  extraction  of  the  metals.  If  the  ore  contains  gold  and  silver 
these  metals  are  to  a  very  large  extent  set  free,  and  are  more  readily 
attacked  by  the  solvent. 



The  objects  of  roasting,  therefore,  may  be  summarized  as  follows: 

1.  To  oxidize.  The  common  elements  oxidized  are  iron,  copper, 
lead,  zinc,  aluminum,  calcium  and  magnesium. 

2.  To  volatilize.  The  common  elements  volatilized  are  sulphur, 
arsenic,  antimony,  bismuth  and  tellurium. 

3.  To  sulphatize.  The  common  elements  sulphatized  are  calcium, 
magnesium  and,  to  some  extent,  lead  and  zinc. 

4.  Dehydration.  The  object  of  dehydration  is  to  agglomerate  the 
ore  particles  and  make  them  more  susceptible  to  leaching,  decantation, 
or  filtration.  Usually  ore,  which  will  percolate  or  filter  very  slowly 
before  roasting,  will  percolate  or  filter  quite  rapidly  after  roasting. 
Mill  dust,  when  raw,  may  be  difficult  to  filter,  but  after  roasting  filtra- 
tion takes  place  quite  rapidly. 

5.  To  make  the  ore  porous.  Oxidation  of  sulphide  and  telluride 
ores,  by  the  elimination  of  the  sulphur  and  tellurium,  must  of  necessity 
make  the  ore  particles  more  porous  and  present  a  greater  surface  to  the 
action  of  the  solvent.  Most  ores  will  yield  a  very  much  better  extraction 
after  roasting  than  before,  even  though  they  are  otherwise  equally 
susceptible  to  treatment. 

6.  To  free  the  gold  and  silver  particles.  Neither  chlorine  nor  cyanide 
are  practical  solvents  of  gold  and  silver  in  their  telluride  combinations, 
and  in  their  sulphide  combinations  they  present  serious  difficulties. 
After  roasting,  the  gold  and  silver  are  in  their  metallic  state  and  are 
readily  soluble  if  the  particles  have  not  been  fused. 

7.  To  convert  the  .desired  metals  into  soluble  form.  Copper,  in  its 
sulphide  combinations,  is  quite  insoluble  in  either  acid  or  alkaline  solu- 
tions, while  in  its  oxide  combinations  it  is  readily  soluble. 

Chemical  Combinations  of  the  Metals  before  Roasting. — Copper,  gold 
and  silver  ores  as  they  come  from  the  mine,  may  contain  any  of  the  base 
elements.  The  matrix  is  almost  always  quartz,  but  associated  with  it 
will  usually  be  found  one  or  more  of  the  elements  enumerated: 

Aluminum;  usually  as  a  silicate,  fluoride  or  sulphate. 

Antimony;  usually  as  a  sulphide. 

Arsenic;  usually  as  a  sulphide. 

Barium;  usually  as  a  sulphate  or  carbonate. 

Calcium;  usually  as  a  carbonate,  fluoride,  or  sulphate. 

Cobalt;  usually  as  a  sulphide. 

Copper;  usually  as  a  sulphide,  carbonate,  oxide  or  silicate. 

Iron;  usually  as  a  sulphide,  carbonate,  or  oxide. 

Lead;  usually  as  a  sulphide  or  carbonate. 

Magnesium;  usually  as  a  carbonate. 

Manganese;  usually  as  an  oxide. 

Nickel;  usually  as  a  sulphide. 

Silver;  usually  as  a  sulphide. 


Zinc;  usually  as  a  sulphide  or  carbonate. 

Sulphur  and  tellurium  are  usually  found  in  combination  with  the 
metals  as  sulphides  and  tellurides. 

General  Chemical  Reactions  During  Roasting.— The  carbonates,  on 
heating,  are  readily  converted  into  the  oxides  of  the  metals  and  carbon 

MC03  =  MO+C02. 
When  metallic  sulphides  are  heated  in  the  presence  of  air,  metallic 
oxides  and  sulphur  dioxide  are  formed: 

MS  +  03  =  MO  +  S02. 
Most  of  the  sulphur  dioxide  passes  off,  but  a  small  portion  of  it  is 
converted  into  sulphur  trioxide  by  contact  with  the  metallic  oxides 
formed,  or  with  the  silica  contained  in  the  ore: 

S02  +  0  +  Si02  =  S03  +  Si02 

Some  of  the  sulphur  trioxide  will  escape,  while  some  will  combine 
with  the  metallic  oxides  to  form  metallic  sulphates : 

MO+S03  =  MS04. 
By  heating,  the  metallic  sulphates  are  dissociated,  some  giving  off 
sulphur  trioxide,  others  sulphur  dioxide  and  oxygen: 
2FeSO,=FeA  +  S03  +  S02 
2CuSO,  =  2CuO  +  2S02  +  02. 
The  sulphates  of  copper,  antimony,  iron,  and  nickel,  are  completely 
decomposed   at   a  red  heat.     A  higher  temperature   decomposes  the 
sulphates  of  aluminum,  silver,  lead,  manganese,  and  zinc.     An  ordinary 
white  heat  has  no  action  on  the  sulphates  of  the  alkalies  and  alkaline 
earths,  potassium,  sodium,  barium,  calcium,  and  magnesium,  but  at  the 
most  intense  heat  procurable,  which  is  never  used  in  a  roasting  furnace, 
the  sulphates  of  barium  and  calcium  are  changed  to  oxides.     At  the 
same  temperature,   sodium   and  potassium  sulphates   are   completely 

Essential  Factors  in  Roasting. — The  essential  factors  in  roasting  are: 

Air,  or  Oxygen. 
These  are  complementary  terms,  and  each  may  be  carried  in  excess, 
to  the  neglect  of  the  others.  Roasting,  as  already  stated,  is  substan- 
tially oxidation  by  the  application  of  heat  and  air.  Mineralized  veins  are 
oxidized  to  great  depths  by  time  and  the  action  of  atmospheric  and 
aqueous  agencies,  without  the  necessity  of  any  perceptible  heat.  In 
this  respect,  it  differs  from  roasting.  If  the  temperature  is  increased  the 
time   of  oxidation  is   diminished.     A  high   temperature,   without   an 


abundance  of  air,  is  of  little  avail.  A  moderate  temperature,  with  an 
abundance  of  air,  is  highly  efficient,  nevertheless  time  is  necessary  to 
effect  complete  oxidation,  and  to  get  the  best  roast  for  subsequent 
chemical  treatment  to  extract  the  metals. 

Time. — Time  in  roasting,  and  in  oxidation,  is  a  most  variable  factor. 
Pyrites  may  be  oxidized  almost  instantly,  in  the  highly  oxidizing  atmos- 
phere of  a  shaft  furnace,  or  it  may  take  countless  ages,  as  in  the  oxidation 
of  mineralized  veins,  where  the  elements  of  both  temperature  and  air 
are  lacking. 

Some  idea  of  the  relation  of  time,  temperature,  and  air  may  be 
obtained  from  roasting  tests  made  in  Denver,  on  Gilpin  County  sulphide 
ore.  The  ore  was  roasted  in  a  three-compartment  shaft  furnace  at  the 
rate  of  75  tons  a  day.  Each  particle  was  exposed  on  all  sides  to  a  highly 
heated  oxidizing  atmosphere.  The  ore  was  thoroughly  roasted  in  13/4 
minutes,  which  was  the  total  time  it  remained  in  the  furnace.  Similar 
ore  was  roasted  in  a  mechanical  reverberatory,  with  a  bed  from  4  to  5  in. 
deep,  and  notwithstanding  that  the  ore  was  rabbled  continuously  and 
remained  in  the  furnace  from  4  to  5  hours,  the  roast  was  not  satisfactory. 
The  difficulty  lay  in  the  inability  of  the  air  to  penetrate  the  deep  bed 
of  ore.  As  soon  as  an  abundance  of  air  was  supplied,  as  for  example, 
when  the  ore  was  discharged  from  the  furnace,  innumerable  sparks 
appeared,  showing  that  the  oxidation  was  taking  place  more  rapidly. 

In  the  Stedefeldt  shaft  furnace  the  ore  is  oxidized  almost  instantly, 
and  the  time  reduced  to  a  minimum,  as  compared  with  roasting  in  the 
ordinary  reverberatory  or  revolving  furnaces.  In  the  Stedefeldt  fur- 
nace, which  was  used  only  for  chloridizing  roasting  silver  ores,  the 
chlorine  acted  as  an  energetic  oxidizer,  and  this  materially  assisted  in 
the  roasting. 

The  best  results  in  roasting  will  usually  be  obtained  when  the  time 
factor  is  made  as  great  as  possible,  and  the  temperature  as  low  as  pos- 
sible, assuming  that  the  air  factor  remains  the  same.  Or,  again,  the 
best  results  will  always  be  obtained  by  having  the  time  and  air  factors 
as  large  as  possible,  speaking,  of  course,  within  practical  limits.  By 
increasing  the  air,  the  temperature  remaining  the  same,  the  time  will  be 
greatly  diminished  without  detriment  to  the  roast. 

If  ore  is  roasted,  as  in  a  shaft  furnace  of  the  Stedefeldt  type,  where 
it  is  showered  through  a  highly  heated  oxidizing  atmosphere,  the  time 
of  roasting  is  reduced  to  a  minimum,  but  the  combustion  of  the  pyrite  is 
likely  to  be  so  intense  that  the  heat  developed  in  the  particle  itself  is 
likely  to  fuse  it.  Careful  panning" of  roasted  sulphide  ore,  roasted  under 
such  conditions,  will  usually  disclose  some  of  the  grains  as  fused  or  even 
shotted,  which  is  the  worst  possible  condition  for  subsequent  treatment 
by  a  solvent  process,  largely  because  of  the  inability  of  the  solvent  to 
penetrate  the  fused  or  shotted  particle. 


Temperature.— The  regulation  of  temperature  in  a  furnace  to  get  the 
best  results  in  roasting,  depends  largely  on  the  nature  and  composition 
of  the  ore.  In  simple  ores,  not  containing  too  much  sulphur,  slow  initial 
heating  is  not  essential  if  care  is  taken  not  to  carry  the  heat  too  near  the 
sintering  point.  Even  on  concentrates,  and  ore  high  in  sulphur,  the 
initial  heat  may  be  reasonably  high  without  deleterious  effects,  provided 
there  is  an  abundance  of  air  and  the  rabbling  in  sufficiently  frequent. 
Copper  sulphides  are  vastly  more  sensitive  to  high  temperatures  than 
iron  sulphides,  and  with  galena,  which  fuses  at  a  low  temperature,  the 
utmost  care  must  be  taken. 

When  roasting  silicious  ore  in  large  furnaces,  or  even  pyritic  material 
containing  only  small  quantities  of  copper  or  lead  sulphides,  the  rear  fire 
may  be  pushed  quite  as  hard. as  the  first  one,  since  the  temperature  of 
the  ore  must  be  brought  to  the  ignition-  or  to  the  volatilization-pomt  of 
sulphur,  before  roasting  can  begin.  This  fact  was  demonstrated  by 
interesting  experiments  made  in  roasting  Cripple  Creek  gold  ore,  in  four 
100-ton  furnaces.  Three  of  these  furnaces  were  of  the  ordinary  me- 
chanical reverberatory  type;  the  fourth  had  a  revolving  hearth,  with  a 
gas  producer  in  the  center,  and  which  was  so  fired  that  the  temperature 
throughout  the  entire  hearth  was  practically  the  same.  The  raw  ore 
entering  the  furnace  was  subjected  to  almost  the  same  heat  as  the 
roasted  ore  being  discharged.  A  comparison  of  several  thousand  tons 
of  tailings  from  the  different  furnaces  showed  no  material  difference  in 
the  extraction. 

The  ordinary  roasting  starts  with  a  low  initial  heat,  and  finishes  at 
the  highest  temperature  the  ore  will  stand  without  sintering.  This  is 
particularly  true  of  all  revolving  furnaces,  and  to  a  large  extent  in  rever- 
beratories  also.  While  this  may  be  the  best  for  some  ores,  careful  com- 
parative tests  in  large  furnaces  would  indicate  that  it  is  better  to  bring 
the  ore  as  quickly  as  possible  to  a  dull  red  (or  even  cherry  red)  heat,  and 
that  the  finishing  temperature  should  not  be  too  high.  In  roasting  ore 
containing  from  2  per  cent,  to  4  per  cent,  sulphur,  in  a  furnace  having 
say  four  fire-boxes  and  roasting  100  tons  per  day,  the  best  results  will  be 
obtained  by  firing  the  finishing  fire  box  at  a  lower  temperature  than  the 
one  preceding  it.  The  dark  magnetic  oxide  is,  to  a  large  extent,  con- 
verted into  the  ferric  oxide  by  the  prolonged  roasting  at  a  moderately 
low  temperature.  Similarly  the  cuprous  oxide,  which  is  with  some 
difficulty  soluble  in  ^cids,  may  be  reduced  to  the  cupric  oxide,  which  is 
quite  readily  soluble. 

If  the  ore  is  overheated,  as  is  frequently  the  case  when  the  finishing 
heat  is  high,  it  will  have  a  dark  appearance;  whereas,  if  finished  at  a 
lower  temperature,  the  ore  will  have  the  red  appearance  of  ferric  oxide. 
Overheating,  or  lack  of  air,  will  convert  the  ferric  oxide  into  the  magnetic 


o^ide,  which,  at  a  lower  temperature  and  with  an  abundance  of  air,  may 
t»e  reconverted  into  ferric  oxide. 

In  order  to  determine  the  effect  of  temperature  on  the  extraction, 
the  following  laboratory  experiments  were  made  on  Cripple  Creek  ore. 
Chlorine  was  used  as  the  solvent.  Head  assay  of  raw  ore,  gold  5.32  oz.; 
head  assay  of  roasted  ore,  gold  5.56  oz.;  sulphur  in  raw  ore,  4.02  per  cent. 

Test  No.  1. — The  ore  was  given  what  appeared  to  be  an  ordinary 
roast  in  an  assay  muffle.  The  finishing  heat  was  that  ordinarily  given 
in  large  furnaces.  The  ore  did  not  show  any  sintering.  Sulphur  in 
roasted  ore,  soluble,  1.16  per  cent.;  insoluble,  0.26  per  cent.;  total,  1.42 
per  cent.  Average  tailing  from  10  bottle  tests,  0.50  oz.  Extraction, 
91  per  cent. 

Test  No.  2. — The  ore  was  given  a  prolonged,  roast  to  reduce  the  sul- 
phur content.  The  finishing  heat  was  quite  high  (about  1575°F.)  but 
the  ore  was  not  sintered.  Sulphur  in  roasted  ore,  soluble,  0.90  per  cent. ; 
insoluble,  0.10  per  cent.;  total,  1.00  per  cent..  Average  tailing  from  10 
bottle  tests,  0.33  oz.     Extraction,  94  per  cent. 

Test  No.  3. — The  ore  was  roasted  at  a  high  heat,  and  finished  at  a 
very  high  temperature  (about  1650  to  1700°  F.) ;  it  had  a  dark  appearance 
and  was  slightly  fused.  Sulphur  in  roasted  ore,  soluble,  0.34  per  cent.; 
insoluble,  0.09  per  cent.;  total,  0.43  per  cent.  Average  tailing  from  10 
bottle  tests,  1.23  oz.  Extraction,  7S  per  cent.  Two  bottles  were 
recharged  and  treated  12  hours;  recharge  tailings  ran  0.79  oz.  Extrac- 
tion, 86  per  cent. 

Test  No.  4. — The  ore  was  roasted  10  hours  in  the  mufHe.  It  was 
brought  quickly  to  a  red  heat,  and  finished  at  a  moderately  high  tem- 
perature. The  ore  had  a  dark  appearance,  and  the  finer  particles  were 
slightly  fused.  Sulphur  in  roasted  ore,  soluble,  0.75  per  cent.;  insolu- 
ble, 0.11  per  cent.;  total,  0.86  per  cent.  Average  tailing  from  10 
bottle  tests,  0.73  oz.     Extraction,  88  per  cent. 

From  Tests  3  and  4  it  is  evident  that  if  the  ore  is  fused  or  sintered, 
a  close  extraction  is  impossible. 

Test  No.  5. — ^The  ore  was  roasted  5  hours  at  a  very  low  temperature 
(scarcely  visible  red) ;  it  was  taken  out  of  the  muffle  and  divided  into 
two  parts.  One  half  was  returned  to  the  muffle  and  roasted  8  hours 
longer  at  a  low  dull  red  heat.  The  finishing  heat  was  a  dull  red.  The 
entire  roasting  was  performed  at  a  prolonged  low  temperature.  Sulphur 
in  roasted  ore,  soluble,  0.74  per  cent.;  insoluble,  0.25  per  cent.;  total,  0.99 
per  cent.  Average  tailing  from  10  bottle  tests,  0.15  oz.  Extraction, 
97.3  per  cent. 

Test  No.  6. — The  other  half  of  the  ore  taken  from  No.  5,  after  roast- 
ing 5  hours,  was  then  put  into  the  muffle  and  roasted  5  hours  more  and 
finished  at  a  higher  temperature.  The  ore  was  not  fused  or  sintered. 
Sulphur  in  roasted  ore,  soluble,  0.69  per  cent.;  insoluble,  0.24  per  cent.; 


total,  0.93  per  cent.  Average  tailing  from  10  bottle  tests,  0.48  oz. 
Extraction,  91  per  cent. 

In  the  bottle  tests  for  the  different  roasts,  the  conditions  were  kept 
the  same.  The  chemicals  corresponded  to  15  and  20  lb.  of  bleach,  and 
30  and  40  lb.  of  sulphuric  acid,  per  ten  of  ore.  The  time  of  treatment, 
on  account  of  the  high  grade  of  the  ore,  was  5  hours.  The  ore  was  ground 
to  16  mesh. 

It  will  be  noticed  that  the  best  results  were  obtained  from  No.  5, 
where  the  finishing  heat  was  quite  low.  By  increasing  the  heat,  the 
extraction  was  not  improved,  as  shown  by  No.  6.  In  No.  1,  the  ore  was 
not  roasted  sufficiently,  as  indicated  by  the  sulphur  analysis  of  1.42 
per  cent.  In  No.  3  the  ore  was  roasted  at  too  high  a  temperature;  the 
sulphur  content  is  low,  0.43  per  cent.  The  sulphur  in  No.  2  represents 
more  nearly  the  mill  roast.  For  ore  having  4  per  cent,  sulphur,  the 
soluble  and  insoluble  sulphur  in  the  roasted  ore,  as  represented  by  No. 
2,  might  be  considered  normal  for  Cripple  Creek  ore.  A  prolonged  low 
heat,  as  represented  by  No.  5,  will  give  the  best  average  extraction. 
The  conditions  there  represented,  however,  could  probably  not  be  fully 
realized  in  practice,  on  account  of  the  reduced  capacity  of  the  furnaces. 

These  test  were  repeated  in  large  furnaces  roasting  100  tons  of  ore 
daily,  and  each  test  represents  a  day's  run,  or  200  tons.  The  extraction 
is  based  on  the  mill  tailings.  Furnace  No.  1  was  fired  with  low  initial 
heat  and  a  higher  finishing  heat.  Furnace  No.  2  was  fired  with  a  higher 
initial  heat  and  a  lower  finishing  heat.  Care  was  taken  to  get  the  best 
possible  roast  under  both  conditions. 

Furnace  Test  No.  1 

Furnace  No.  1  Furnace  No.  2 

Sulphur,  raw  ore,  2 .  75  per  cent.  2 .  50  per  cent. 

f  soluble,  0 .  75  per  cent.  0 .  79  per  cent. 

Sulphur  in  roasted  ore,       I   insoluble,  0 .  09  per  cent.  0 .  14  per  cent. 

[  total,  0 .  84  per  cent.  0 .  93  per  cent. 

Assay  of  roasted  ore,  gold,  1 .05  oz.  1 .85  oz. 

Assay  of  chlorination  tailings,  0.08  oz.  0. 11  oz. 

Extraction,  92 . 4    per  cent.  94 . 6    per  cent. 

Furnace  Test  No.  2 

Sulphur,  raw  ore,  2 .  66  per  cent.  2 .  66  per  cent. 

f  soluble,  0.71  per  cent.  0.77  per  cent. 

Sulphur,  roasted  ore,      <   insoluble,  0 .  07  per  cent.  0 .  08  per  cent. 

[  total,  0 .  78  per  cent.  0 .  85  per  cent. 

Assay  of  roasted  ore,  gold,  1 .  23  oz.  1.11  oz. 

Assay  of  chlorination  tailings,  0.12  oz.  0.07  oz. 

Extraction,  90.0     percent.  94.0    percent. 

These  comparative  tests  in  actual  mill  practice,  in  furnaces  roasting 
100  tons  of  ore  per  day,  are  characteristic  of  many  others  made  along 



the  same  linea.  The  results  clearly  indicate  that  the  lower  finishing 
heat  gives  the  best  average  results;  that  ore  roasted  at  a  low  temperature 
gives  up  its  values  better  than  ore  roasted  at  a  high  temperature,  and 
that  sintering  or  overheating  is  highly  injurious.  The  ore  in  these  tests 
was  ground  to  10  mesh.  Time  of  chlorination,  3  hours.  The  chemicals 
used  were  15  lb.  of  bleach  and  30  lb.  of  sulphuric  acid,  per  ton  of  ore. 
Valentine's  Temperature  Experiments. — Valentine'  made  interesting 
experiments  on  the  effect  of  temperatures  on  iron  pyrite,  with  and  with- 
out free  access  of  air.  The  results  of  his  experiments  undertaken  to 
ascertain  the  effect  of  heat  on  FeSj,  when  air  is  freely  given  access,  are 
given  as  follows : 


Duration  of 

S.  in  residue,   j 


Per  cent,  of 

temperature,  deg.  F. 


per  cent. 


per  cent. 

S.  expelled 

Original  pyrite 

53.43           ' 

1  hour 





2  3/4  hours 





20  minutes 



!       98.54 


45  minutes 





20  minutes 





1  hour 





15  minutes 





20  minutes 





35  minutes 



:       97.10 


2  hours 




It  will  be  noticed  that  a  larger  amount  of  sulphur  remains  in  the 
residues  when  higher  temperatures  have  been  applied. 

Valentine,  from  his  experiments,  draws  the  following  conclusions 
in  roasting  pyritic  ores: 

1.  Heat  alone  without  access  of  air,  can  remove  at  best  only  one 
half  of  the  sulphur  present. 

2.  Atmospheric  oxygen  is  absolutely  necessary  for  a  proper  desul- 

3.  Even  at  a  low  heat,  ore  is  properly  desulphurized  if  air  can  gain 
access  freely  to  the  FeSj  in  it. 

4.  Sulphate  of  iron  can  be  decomposed  equally  well  with  or  with- 
out air. 

.5.  In  order  that  the  residuum  sulphur  in  roasted  ore  may  consist  as 
far  as  possible  of  sulphates,  the  roasting  must  be  done  under  free  access 
of  air. 

'  Trans.  A.  I.  M.  E.,  Vol.  XVIII. 


6.  Fusion  or  sintering  of  ore  is  likely  to  retard  further  desulphur- 

7.  Sintering  does  not  allow  much  of  the  remaining  sulphur  to  be  in 
the  form  of  sulphate. 

Air. — Much  oxygen  is  consumed  in  roasting  sulphides,  and  a  highly 
oxidizing  atmosphere  is  essential  to  good  results.  With  the  sulphur 
fumes,  and  the  products  of  combustion  from  the  fire-boxes  passing  over 
the  partly  roasted  ore  in  the  rear  of  the  furnace,  the  atmosphere,  while 
it  may  not  be  strongly  reducing,  is  certainly  not  highly  oxidizing.  A 
comparatively  small  amount  of  sulphur  dioxide  in  the  furnace  gases 
will  greatly  retard  oxidation;  and  if,  in  addition  to  sulphur  dioxide,  the 
atmosphere  is  charged  with  carbon  dioxide  from  the  combustion  of  the 
fuel,  effective  roasting  is  impossible.  The  only  advantage  to  be  gained 
in  passing  these  deleterious  gases  over  the  fresh  ore  as  it  is  introduced 
into  the  roasting  furnace  is  to  heat  it  so  that  oxidation  can  proceed  more 
rapidly  when  it  reaches  a  more  highly  oxidizing  atmosphere. 

It  is  desirable  to  bring  the  sulphides  to  the  ignition  temperature  as 
soon  as  possible  after  the  ore  has  been  introduced  into  the  furnace. 
The  value  of  the  vitiated  hot  gasses  passing  over  a  long  stretch  of  cold 
ore  to  heat  it  and  thereby  save  fuel  is  largely  overestimated.  The  loss 
will  exceed  the  gain. 

With  a  deep  bed  of  ore  in  the  furnace,  say  from  3  1/2  to  .5  in.,  even 
in  a  highly  oxidizing  atmosphere,  only  the  ore  on  the  surface  is  under 
thorough  oxidizing  conditions,  while  that  below  the  surface  is  not  so 
advantageously  placed.  If  overheated,  therefore,  fusion  or  matting  is 
likely  to  occur  in  the  early  stages  of  the  roasting,  and  when  the  ore  is 
heated  too  quickly  to  get  a  correspondingly  quick  oxidation.  If  fusion 
occurs,  the  particles  assume  a  dark,  and  sometimes  glazed,  appearance. 
In  this  condition  it  is  more  difficult  to  sufficiently  eliminate  the  remaining 
sulphur.  The  metals,  too,  are  difficult  to  extract  in  the  chemical  process, 
owing  to  the  inability  of  the  solvent  to  penetrate  the  pores  of  the  ore 

Speaking  within  practical  limits,  it  is  not  so  much  the  high  tempera- 
ture as  the  lack  of  air  that  is  fatal  to  rapid  and  thorough  roastino-  in 
reverberatory  furnaces.  Fusion  of  the  sulphide  particles  invariably 
occurs  when  the  ore  is  brought  too  suddenly  against  a  high  temperature 
with  insufficient  air.  The  tendency  is  to  convert  the  sulphide  into 
matte.  Much  of  the  sulphur  in  the  deeper  portions  of  the  bed  vola- 
tilizes as  such,  and  when  it  reaches  the  surface  it  burns  to  sulphur  dioxide. 
Heat,  without  access  of  air,  can  remove  only  about  50  per  cent,  of  the 
sulphur  originally  in  the  ore. 

In  order  to  determine  the  effects  of  time  and  air  on  roasting  and 
extraction,  comparative  tests  were  made  with  furnaces  roasting  100  tons 
of  Cripple  Creek  ore  daily,  containing  about  2.75  per  cent,  sulphur. 



The  ore  in  furnace  No.  1  was  roasted  under  normal  conditions;  the  bed 
of  ore  was  about  2  1/2  in.  deep;  the  angle  of  the  rabble  blades  was  22  1/2 
degrees,  and  the  ore  was  about  2  1/2  hours  in  passing  through.  In 
furnace  No.  2  the  angle  of  the  rabble  blades  was  changed  to  12  degrees, 
which  resulted  in  having  the  bed  about  4  1/2  in.  deep;  the  ore  remained 
in  the  furnace  about  5  hours  to  get  the  same  capacity.  All  the  other 
conditions  remained  the  same,  so  far  as  they  could  be  kept  the  same. 

Test  No.  1 

Sulphur,  raw  ore, 

I  soluble, 
Assay,  roasted  ore,  gold. 
Assay,  chlorination  tailings, 

Sulphur,  raw  ore, 

r  soluble, 
Sulphur,  roasted  ore   <    nsoluble 

[  tota  , 
As.say,  roasted  ore,  gold. 
Assay,  chlorination  tailings. 

Sulphur,  raw  ore, 

Assay,  roasted  ore,  gold, 
Assay,  chlorination  tailings. 

Furnace  No.  1 
90  tons  per  day 

2 .  63  per  cent. 

0 .  73  per  cent. 

0.17  per  cent. 

0 .  90  per  cent. 

1.07  oz. 

0 .  09  oz. 
91.5    percent. 

Test  No.  2 

Furnace  No.  1 
100     ons  per  day 
2. 80  per  cent. 
0.79  per  cent. 
0. 14  per  cent, 
0.93  per  cent. 
1.80  oz. 
0.12  oz. 
93 . 3    per  cent. 

Test  No.  3 

Furnace      No.  1 
100  tons  per  day 

2 .  62  per  cent. 

0.76  per  cent. 

0 .  14  per  cent. 

0 .  90  per  cent. 

1 .  90  oz. 
0.12  oz. 

94 . 0   per  cent. 

Furnace  No.  2 

90  tons  per  day 

2.53  per  cent. 

0.94  per  cent. 

0 .  26  per  cent. 

1 .  20  per  cent. 
2.22  oz. 
0.28  oz. 

87 . 4    per  cent. 

Furnace  No.  2 
70  tons  per  day 

2 .  75  per  cent. 

0 .  72  per  cent. 

0.25  per  cent. 

0.97  per  cent. 

1.19  oz. 

0.19  oz. 
85 . 3    per  cent. 

Furnace  No.  2 
70  tons  per  day 

2 .  50  per  cent. 

0. 59  per  cent. 

0 .  23  per  cent. 

0 .  82  per  cent. 

1.08  oz. 

0.29  oz. 
73 . 0    per  cent. 

In  f.urnace  No.  2  it  was  soon  found  that  a  capacity  of  100  tons  per 
day  was  out  of  the  question.  The  capacity  was  at  once  reduced  to  70 
tons  to  give,  what  at  least  appeared  to  be,  a  fair  roast.  The  bed  of  ore, 
which  with  a  capacity  of  100  tons,  was  about  5  in.  deep,  with  70  tons, 
was  reduced  to  4  in.  The  high  tailings  in  No.  2  may  have  been  due,  in  a 
measure,  to  the  higher  temperature  frequently  necessary  to  eliminate 
the  sparks  from  the  roasted  ore. 


It  will  also  be  seen  from  the  sulphur  analyses  that  whilein  test  No.  1, 
the  sulphur  is  higher  in  furnace  No.  2  than  in  furnace  No.  1,  in  tests  2 
and  3 ,  it  is  lower ;  nevertheless  the,  extraction  was  not  improved.  It  will  be 
noticed,  however,  that  the  insoluble  sulphur  in  furnace  No.  2  is  abnormally- 
high  as  compared  with  the  insoluble  sulphur  in  furnace  No.  1.  This  is 
evidently  due  to  lack  of  air,  and  perhaps  higher  temperature,  in  furnace 
No.  2  to  get  approximately  the  same  total  sulphur  elimination  as  in  furnace 
No.  1. 

By  increasing  the  bed  from  2  or  2  1/2  in.  to  4  or  5  in.  which  makes 
the  penetration  of  the  air  more  difficult,  the  capacity  was  reduced  from 
100  to  70  tons  per  day,  and  the  quality  of  the  roast  was  very  inferior, 
notwithstanding  that  the  time  of  roasting  was  practically  doubled. 

The  ore  in  these  tests  was  crushed  to  12  mesh,  and  chlorinated  3 
hours  with  a  chemical  charge  of  15  lb.  of  bleach  and  20  lb.  of  acid  per 
ton  of  ore. 

The  time  of  roasting,  of  5  hours,  in  reverberatory  furnace  No.  2, 
may  be  compared  to  the  time  the  ore  is  subjected  to  roasting  in  a  shaft 
furnace,  which  may  be  considered  about  half  a  minute ;  or  1  /  600  of  the  time. 
The  temperature  in  both  cases  may  be  considered  the  same;  the  difference 
in  the  results,  therefore,  is  due  to  the  difference  in  air  supply. 

Interesting  experiments  were  made  in  Denver  to  determine  the  effect 
of  an  abundance  of  air  supply  in  roasting  charges  of  2000  lb.  of  ore  in 
a  hand-rabbled  reverberatory  furnace.  In  these  experiments,  some  of 
the  charges  were  roasted  in  the  ordinary  way,  while  in  others  arrange- 
ment was  made  to  pass  air  through  the  incandescent  roasting  ore,  both 
by  up-draft  and  down-draft;  other  conditions  remained  the  same.  The 
experiments  proved  that  the  capacity  of  the  furnace,  due  to  the  extra 
air  supply,  was  trebled  in  roasting  a  heavy  sulphide  ore;  an  appreciable 
saving  of  fuel  was  effected,  and  the  sulphur  elimination  was  more  perfect. 

The  amount  of  air  required  in  practice  in  roasting  is  enormously  in 
excess  of  that  required  to  combine  with  the  sulphur  and  other  elements. 
Theoretically,  at  least,  the  air  in  all  parts  of  the  furnace  should  be  kept 
as  pure  as  possible;  on  the  other  hand,  the  cost  of  heating  a  large  volume 
of  excess  air  is  considerable.  In  practice,  therefore,  the  best  results  will 
be  obtained  by  carefully  balancing  these  two  opposing  factors. 

The  amount  of  sulphur  dioxide  in  the  flue  gases  for  effective  roast- 
ing should  not  exceed  2  per  cent.  When  the  sulphur  dioxide  in  the 
furnace  atmosphere  reaches  4  per  cent.,  roasting  becomes  slow;  when 
it  reaches  8  per  cent.,  it  becomes  very  slow;  and  when  it  reaches  9  per 
cent,  and  over,  the  reactions  practically  cease. 

Rabbling. — Rabbling  is  an  important  operation  in  roasting.  Its 
object  is  essentially  to  expose  fresh  particles  of  the  ore  to  the  direct  action 
of  the  air  and  heat,  and  to  facilitate  bringing  the  entire  mass  of  ore  to 
incandescence,  and  thus  assuring  a  uniform  roast. 


It  is  evident  that  the  roasting  is  facilitated  by  frequent  rabbling, 
but  the  frequency  of  the  rabbling  is  limited  by  the  rabbling  mechanism 
in  mechanical  furnaces,  and  by  the  fatigue  of  the  roasterman,  in  hand- 
rabbled  furnaces.  Theoretically,  the  more  the  ore  is  rabbled,  the  better 
will  be  the  roast,  and  this  theoretical  condition  should  be  approached  as 
closely  as  possible.  It  is  for  this  reason,  more  than  all  others,  that 
mechanical  furnaces  give  a  much  better  roasted  product  than  hand- 
rabbled  furnaces.  No  hand  rabbling,  on  a  large  scale,  can  approach 
the  frequency  and  uniformity  of  mechanical  rabbling. 

Effect'  of  Metallic  Sulphides  if  Heated  with  Exclusion  of  Air.— Gold 
and  platinum  can  be  completely  desulphurized.  The  sulphide  of  silver 
(AgS)  remains  undecomposed.  The  sulphides  of  arsenic,  antimony, 
and  mercury,  volatilize  unchanged.  Iron  pyrites  (FeSj)  gives  up  23 
per  cent,  of  its  sulphur,  and  is  reduced  to  magnetic  pyrites  (FegSg),  which 
by  a  strong  heat  may  be  reduced  to  ferrous  sulphide  (FeS).  The  ferrous 
sulphide  is  not  further  reducible.  Of  the  copper  minerals,  chalcocite  (CujS) 
is  not  decomposed,  but  the  chalcopyrite  (CuTeS,)  loses  only  one  part  of 
the  sulphur  which  is  combined  with  the  iron.  Galena  (PbS)  is  reduced 
to  a  lower  stage  with  separation  of  metallic  lead. 

Sulphur. — Sulphur  usually  occurs  combined  with  the  base  metals  as 
sulphide,  but  not  infrequently  the  ore  is  highly  charged  with  sulphates, 
due  to  partial  decomposition  by  atmospheric  and  aqueous  agencies. 
Sulphur,  combined  with  some  of  the  metals  as  sulphide  or  sulphate,  is 
highly  injurious  in  the  hydrometallurgical  process;  and  if  occurring  in 
large  quantities,  it  is  fatal.  Many  of  the  sulphates  are  acted  upon  by 
acids;  in  any  cases  the  soluble  sulphates  affect  the  leaching  solution 
injuriously.  In  the  cyanide  process  some  of  the  sulphides,  as  for  example 
pyrite,  are  not  particularly  injurious,  while  most  of  the  sulphates  offer 
difficulties.  In  the  chlorination  process,  the  sulphur  in  combination 
with  some  of  the  metals  is  displaced  by  the  chlorine,  which  itself  unites 
with  the  metal  or  acts  as  an  oxidizer.  In  either  case,  the  chemicals  are 
consumed  by  reacting  with  the  base  elements,  and  are  not  available  for 
action  on  the  desired  metals.  Roasting,  in  any  event,  largely  overcomes 
these  difficulties,  and  in  many  cases  practically  eliminates  them  entirely. 

Sulphur  is  rarely,  if  ever,  entirely  eliminated  during  the  roasting. 
Frequently  that  which  remains  is  not  injurious  to  the  process.  The 
sulphur,  as  sulphide  or  sulphate,  may  be  encased  in  quartz  particles,  or 
it  may  be  in  the  ore  as  sulphates  of  the  alkali  metals  or  of  the  alkaline 
earths.  So  combined,  it  is  not  replacable  by  any  of  the  chemicals  ordi- 
narily used.  The  sulphates  of  sodium,  potassium,  barium,  calcium,  and 
magnesium  appear  to  be  unaffected  by  either  hydrochloric  and  sulphuric 
acids,  chlorine,  bromine,  cyanide,  or  sodium  hyposulphite.  Their  pres- 
ence, in  the  solution,  may  however  have  some  effect  on  the  solubilities 
of  the  various  solvents. 


Most  of  the  sulphur  in  either  copper,  gold  or  silver  ores  is  usually 
combined  with  iron,  as  iron  pyrite  (FeS^) .  One  of  these  atoms  of  sulphur 
may  be  distilled,  or  be  burned  to  sulphur  dioxide  at  a  low  temperature. 
In  the  roasting  of  sulphides,  sulphur  dioxide  is  exclusively  formed.  In 
the  presence  of  air,  by  catalytic  action  with  indifferent  substances  such 
as  silica  or  iron  oxide,  there  is  always  formed  a  small  amount  of  sulphur 
trioxide,  which  with  the  moisture  of  the  air  and  that  contained  in  the  ore 
and  fuel  gases,  gives  sulphuric  acid. 

The  elimination  of  sulphur  from  concentrates  or  heavy  sulphide 
ore  is  accompanied  by  the  evolution  of  considerable  heat.  Concentrates 
containing  from  25  to  35  per  cent,  sulphur  are  frequently  roasted  down  to 
5  per  cent,  by  the  heat  generated  from  their  own  oxidation.  Thirty-two 
parts  0^  sulphur,  in  combining  with  32  parts  of  oxygen  (that  is,  forming 
SO2),  evolves  69,260  heat  units;  and  if  the  oxidation  proceeds  to  SO3, 
91 ,900  heat  units  are  evolved.  These  figures  may  be  compared  with  those 
which  correspond  to  the  passage  of  carbon  into  carbon  monoxide  (CO) 
and  carbon  dioxide  (COJ  when  29,160  and  97,200  units  of  heat,  respec- 
tively, are  evolved.  The  evolution  of  heat  by  the  rapid  oxidation  of 
sulphur  is  practically  demonstrated  in  the  various  sinter-roasting 
processes  in  which  copper  and  lead  sulphide  ore  and  fines  are  fused  into 
a  coherent  mass  by  the  heat  from  the  sulphur  alone. 

The  elimination  of  sulphur,  in  roasting,  varies  greatly  with  different 
ores.  Some  forms  of  pyrite  are  more  difficult  to  roast  than  others. 
Unoxidized  ores  from  the  deeper  levels  of  a  mine  are  more  difficult  to 
roast  than  the  partially  oxidized  ores  nearer  the  surface,  even  though 
the  sulphur  content  of  both  is  approximately  the  same.  The  chemical 
composition  of  the  ore,  aside  from  its  sulphur  content,  has  much  to  do 
with  the  roasting.  Ferrous  sulphate,  for  example,  is  much  more  easily 
broken  up  than  zinc  sulphate,  or  than  the  sulphates  of  the  alkalies  or 
alkaline  earths.  Ore  containing  much  lime  is  likely  to  be  high  in  sulphur 
after  roasting. 

In  some  of  the  Cripple  Creek  ores  having  2.75  per  cent,  sulphur,  the 
best  extraction  is  frequently  made,  and  without  undue  consumption  of 
chemicals,  when  the  roasted  ore  contains  from  0.60  to  0.80  per  cent, 
sulphur.  When  the  total  sulphur  in  the  roasted  ore  is  less  than  0.40  per 
cent,  the  tailing  are  usually  high.  Other  Cripple  Creek  ores,  which  are 
partially  oxidized,  give  the  best  extraction  when  the  sulphur  is  from  0.30 
to  0.50  per  cent.,  the  insolubles  usually  going  from  0.03  to  0.08  per  cent., 
and  the  solubles  from  0.35  to  0.40  per  cent. 

It  is  customary  in  many  of  the  mills  to  make  frequent  sulphur  deter- 
minations. Sometimes  they  are  made  for  such  shift  for  every  furnace; 
sometimes  once  a  day.  These  sulphur  determinations  are  made  both 
for  soluble  and  insoluble  sulphur.  The  insoluble  sulphur  is  more  par- 
ticularly relied  upon  to  indicate  the  roast.     The  soluble  sulphur  is  that 


which  is  soluble  in  boiling  water;  usually  a  little  sodium  carbonate  is 
added  before  boiling. 

The  progress  in  the  elinaination  of  the  sulphur,  when  treating  100 
tons  of  Cripple  Creek  ore  daily  in  large  mechanical  furnaces,  is  shown 
by  the  following  samples  taken  at  various  points  in  the  furnace  during  the 
roasting.  The  ore  remained  in  the  furnace  about  2  3/4  hours,  so  that 
the  distance,  in  feet,  from  the  feed  will  also  closely  represent  the  time 
in  minutes  for  the  ore  in  the  furnace,  when  the  respective  samples  were 
taken.  The  ore  was  rabbled  every  17  seconds.  The  results  are  averages 
of  a  large  number  of  sets  of  samples  taken  from  three  different  furnaces. 
The  samples  were  taken  so  as  to  fairly  represent  the  total  cross  section  of 
the  ore.  No.  3  was  taken  from  a  type  of  furnace  totally  different  from 
the  others.     The  ore  before  roasting  was  crushed  to  12  mesh. 

No.  3  furnace  had  a  revolving  hearth  and  was  fired  at  a  lower  tem- 
perature than  the  others.  Notwithstanding  the  high  sulphur  content 
in  the  roasted  ore  from  No.  3,  the  extraction  by  chlorination  was  about 
the  same  as  for  the  others.  The  tailings  from  No.  2  were  somewhat 
higher  than  from  No.  1;  this  was  doubtless  due  to  the  fact  that  in  No.  2, 
with  three  fire-boxes,  the  ore  had  to  be  roasted  at  a  higher  temperature 
than  in  No.  1,  which  had  four  fire-boxes. 




Furnace  No.  1   (Four  Fire-boxes) 

Sample  taken,  feet  (also 
approximate  time  in 
min.)  from  feed 

Raw  ore 

45  ft.  Minutes. . 
65  ft.  Minutes. . 
100  ft.  Minutes. , 
160  ft.  Minutes. . 


of  fire-box 

from  feed 

in  feet 





per  cent. 


per  cent. 



per  cent. 

and  differ- 
ence in  per 

Per  cent 
of  S  com- 
pared to 
raw  ore 

Per  cent 
of  S  com- 
pared to  S 
























Furnace  No.  2  (Three  Fire-boxes) 





0.85  ;45 
0.59  ;31 
0.29  ;  15 
0.16  ;    8 



65  ft.     Minutes 

100  ft.    Minutes 

160  ft.    Minutes 







Furnace  No.  3  (Revolving  Hearth) 
















165  min.  from  teed  (roasted  ore) 


In  these  results  it  will  be  noted  that  much  of  the  sulphur  was  driven 
off  early  in  the  operation,  and  before  passing  the  first  fire-box.  This 
is  a  practical  demonstration  of  the  instability  of  the  first  atom  of  sulphur 
in  iron  pyrite.  The  difficulty  of  eliminating  the  last  25  per  cent,  is 
apparent;  the  difficulty  of  eliminating  the  third  25  per  cent,  is  considerable. 
Only  a  small  fraction  of  the  total  sulphur  is  expelled  during  the  last  100 
minutes  of  the  160  minutes  of  roasting,  while  a  large  portion  (about  one- 
fourth)  still  remained  in  the  ore.  Roasting  a  low-sulphur  ore  down  to  a 
trace  is  evidently  as  difficult  as  to  extract  all  but  a  trace  of  the  met- 
als. Roasting  sulphide  ores  down  to  a  "trace  of  sulphur"  and  extract- 
ing all  but  a  "trace  of  the  metals"  are  operations  frequently  spoken  of 
but  rarely  truthfully  realized. 

It  will  be  noticed  that  only  from  60  to  70  per  cent,  of  the  sulphur 
was  eliminated  in  these  roasts.  It  is  safe  to  say  that  fully  two-thirds 
of  the  total  fuel  was  consumed  in  expelling  only  a  small  fraction  of  a 
per  cent,  during  the  latter  half  of  the  operation.  In  the  ordinary 
reverberatory  furnace  there  does  not  appear  to  be  any  adequate  gain, 
in  the  latter  part  of  the  roasting,  for  the  fuel  expended. 

The  tables  of  the  progress  of  roasting  give  a  fair  idea  as  to  the  rate  of 
decrease  of  the  insoluble  sulphur  and  increase  of  the  soluble  sulphur. 
After  the  first  65  or  70  minutes,  the  principal  result  accomplished  by 
roasting,  is  the  changing  of  the  remaining  insoluble  sulphur  to  the 
soluble.  In  furnace  No.  3,  for  example,  there  is  only  a  difference  of  0.11 
per  cent,  in  the  total  sulpur,  between  the  roasted  ore  and  the  first  70 
minutes  of  roasting,  but  for  the  remaining  85  minutes,  0.30  per  cent, 
insoluble  sulphur  was  changed  to  0.10  per  cent.,  and  this  represents  the 
difference  between  a  go.od  and  a  poor  roast. 

It  sometimes  happens  that  the  sulphur  in  the  roasted  ore  from  the 
cooler  is  higher  than  the  discharge  from  the  furnace;  this  may  be  ac- 
counted for  by  the  fact  that  the  rabbles  frequently  push  more  or  less 
partly  roasted  ore  ahead  of  them  in  the  grooves  made  by  the  preceding 
rabble,  or  the  rabbles  themselves  may  carry  partly  roasted  ore  through 
the  furnace  and  discharge  it  on  the  cooler. 

Within  certain  limits,  the  sulphur  content  of  the  roasted  ore  does  not 
appear  to  affect  the  extraction;  beyond  these  limits  the  effect  is  marked. 
Nothing  would  be  gained  in  extraction  by  roasting  the  ore  represented 
in  the  tables  to  say  0.05  per  cent,  insoluble  and  0.20  per  cent,  to  0.40  per 


cent,  soluble  sulphur,  while  the  extra  cost  of  roasting  to  such  a  low  sul- 
phur content  would  be  enormously  increased.  There  would  also  be 
high  tailings  from  overheating. 

There  is  quite  as  much  danger  of  over-roasting  as  under-roasting. 
Ore  roasted  too  much  will  give  high  tailings,  nor  can  these  tailings  be 
materially  reduced  by  repeated  charges  of  chemicals.  If  ore  is  under- 
roasted,  repeated  charges  of  chemicals  may  be  necessary  to  get  the 
desired  extraction;  but  the  tailings  will  be  reduced  each  time,  and  ulti- 
mately the  limit  of  extraction  may  be  obtained. 

An  experiment  was  made  with  a  100-ton  furnace  to  determine  the 
effect  of  ultimate  roasting  on  the  extraction.  The  furnace  was  fired 
under  normal  conditions;  but  instead  of  treating  the  ore  after  its  first 
passage  through  the  furnace,  it  was  returned  again  and  again  for  12  hours. 
It  is  needless  to  say  that  the  ore  was  roasted  "dead";  nevertheless  four 
charges  of  chemicals  on  this  ore  failed  to  give  even  the  average  extraction. 

One  of  the  essential  features  of  roasting  is  to  find  the  sulphur  deter- 
minations which  will  give  the  best  results,  and  to  find  the  point  where  a 
lower  sulphur  content  will  not  appreciably  increase  the  extraction. 
The  sulphur  in  the  roasted  ore,  from  one  mine  or  from  one  district,  which 
has  proved  to  give  the  best  results,  might  be  fatal  to  the  treatriient  of 
ore  from  another  district.  This  is  largely  due  to  the  way  in  which  the 
sulphur  is  combined.  Soluble  sulphur,  as  sodium  or  potassium  sulphate, 
is  unaffected  by  the  chemical  solvents;  if  the  same  amount  of  sulphur 
were  combined  with  iron,  as  ferrous  sulphate,  the  roast  would  be  abso- 
lutely worthless.  Again,  the  insoluble  sulphur,  as  barium  or  calcium 
sulphate,  is  not  particularly  detrimental  to  the  subsequent  treatment; 
but  if  the  same  amount  of  sulphur  is  combined  as  sulphide,  it  is  almost  sure 
to  be  fatal.  Insoluble  sulphur  does  not  necessarily  imply  that  the 
sulphur  is  in  the  form  of  sulphide. 

In  the  roasts,  as  shown  in  the  accompanying  tables,  the  iron  sul- 
phate was  practically  eliminated  at  100  ft.  from  the  feed,  and  totally 
eliminated  ait  120  ft.  In  some  instances  it  was  totally  eliminated  at 
100  ft.,  as  shown  by  the  ferricyanide  test,  which  did  not  produce  the 
usual  delicate  reaction  for  iron.  The  ferricyanide  test,  except  for  making 
a  rough  determination,  is  absolutely  worthless,  since  it  shows  only  the 
sulphur  combined  with  the  iron  as  soluble  sulphate.  It  frequently 
happens  that  the  ore  is  far  from  being  roasted,  when  the  sulphate  of  iron 
is  all  decomposed. 

The  barium  chloride  test,  except  as  a  rough  indication,  is  also  worth- 
less, since  it  precipitates  sulphur  that  might  be  considered  as  perfectly 
harmless.  If  the  ore,  as  shown  in  the  tables,  had  been  roasted  so 
thoroughly  that  no  sulphur  had  been  precipitated  with  barium  chloride, 
it  is  safe  to  say  that  only  a  comparatively  few  tons  of  ore  could  be  roasted 
in  a  day,  and  the  tailings  would  be  quite  sure  to  be  high.     A  considerable 


temperature  would  be  required  to  break  up  the  sulphates  of  the  alkaline 
earths,  and  such  a  temperature  would  be  detrimental  to  the  extraction  of 
the  desired  metals. 

A  direct  sulphur  determination  seems  to  be  the  only  way  of  indicating 
the  roast  with  any  degree  of  accuracy.  It  is  not  so  much  a  matter  of 
absolute  refinement  in  the  sulphur  determination  as  long  as  the  results 
are  uniform.  Nevertheless,  care  and  accuracy  are  essential  to  uni- 
formity. The  essential  idea  of  the  sulphur  determination  is  to  indicate 
the  quality  of  the  roast  as  compared  with  the  extraction.  It  is  not  very 
material  whether  the  sulphur  is  relatively  high  or  low.  If  a  certain 
sulphur  determination  indicates  a  good  or  a  bad  roast  one  day,  theoretic- 
ally at  least,  it  should  indicate  the  same  at  any  other  time.  It  is  evident 
that  these  ideal  conditions  cannot  always  be  realized,  because  the  work- 
ing conditions  of  the  furnace  change  from  time  to  time.  For  example, 
ore  which  is  over-roasted  for  four  hours  on  a  shift  and  under-roasted 
the  other  four,  if  averaged,  might  give  the  same  sulphur  determination 
as  ore  which  had  been  evenly  and  uniformly  roasted  for  the  total  eight 
hours;  but  the  tailings  would  be  entirely  in  favor  of  the  latter  roast. 
Unless  it  is  known  that  the  conditions  of  operating  the  furnaces  have 
changed,  it  is  fair  to  assume  that  they  have  remained  the  same,  and  this 
is  usually  the  case  in  well  conducted  plants. 

In  order  to  get  uniformity  in  the  roasting  operation  and  in  the 
roasted  product,  it  is  essential  that  the  raw  ore  fed  into  the  furnace 
should  average  about  the  same  in  sulphur.  If  the  ore  comes  from 
different  mines,  or  from  different  levels  of  the  same  mine,  it  should  be 
mixed.  Extreme  care  in  mixing  is  neither  necessary  nor  profitable. 
It  is  more  economical  to  build  the  furnace  of  ample  capacitj^,  so  that 
small  variations  in  the  ore  will  naturally  be  taken  care  of,  under  uniform 
conditions  of  operation. 

The  quantity  of  sulphur  which  roasted  ore  may  contain  without 
particular  detriment  is  variable,  depending  largely  upon  the  lime,  mag- 
nesia, and  lead. 

Careful  tests  have  demonstrated  that  an  abundance  of  air  is  not  con- 
ducive to  the  formation  of  sulphates.  Air  does  not  appear  to  be  neces- 
sary to  decompose  sulphates;  nevertheless,  when  it  is  supplied  in  abun- 
dance the  decomposition  of  the  sulphates  is  greatly  facilitated.  -Other 
substances,  by  catalytic  action,  may  also  aid  in  their  decomposition. 

Decomposition  Temperature  of  the  Various  Sulphates. — With  the  ex- 
ception of  lead  sulphate,  all  the  common  metallic  sulphates  are  com- 
pletely decomposed  upon  heating,  into  metallic  oxide,  sulphur  trioxide, 
sulphur  dioxide,  and  oxygen.  Some  give  up  their  trioxide  readily  at 
low  temperatures,  others  require  considerable  heat  and  much  time,  to 
be  completely  freed  from  sulphur.  Kerl,  in  1881,  arranged  the  prin- 
ciple metallic  sulphates,  as  they  are  decomposed  by  a  rising  temper- 



ature,  in  the  following  order:  silver,  iron,  copper,  zinc,  nickel,  cobalt, 
manganese,  and  lead.  Lead  sulphate  is  decomposed  only  at  a  white 
heat.  Bradford'  found  that  ferrous  sulphate  is  decomposed  at  590°  C, 
cupric  sulphate  at  653°  C,  and  silver  sulphate  at  1095°  C.  H.  0. 
Hofman-  found  that  zinc  sulphate  is  decomposed  at  739°  C. 

In  the  presence  of  air  and  other  gases,  and  various  other  substances 
in  the  ore,  the  temperature  of  the  decomposition  of  the  various  metallic 
sulphates  may  be  vitally  affected.  The  decomposition  of  silver  sulphate 
takes  place  at  from  860  to  870°  C.  in  the  presence  of  cupric  oxide,  silica, 
and  ferric  oxide.  In  the  presence  of  reducing  gases,  silver  sulphate 
is  decomposed  at  a  very  moderate  heat,  resulting  in  the  formation  of 
rnetallic  silver. 

Prof.  H.  O.  Hofman  and  W.  Wanjukow  determined  the  decomposition 
temperature  of  some  metallic  sulphates  in  a  current  of  air,  as  follows: 



















Deg.  C. 

Deg.  F. 

150                       302 


530           j             986 


540           1            1004 


653           '           1207 











718           I            1324 


739                       1362 


830                     1526 


850                     1562 


875                     1607 








(New  York  meeting  of  the  American  Institute  of  Mining  Engineers,  Feb.,  1912.) 

Amount  of -Sulphur  Trioxide,  (SO3)  in  the  Sulphur  Dioxide  (SOj) 
Escaping  from  Roasting  Furnaces. — In  the  roasting  of  Spanish  py- 
rites for  the  manufacture  of  sulphuric  acid  it  was  found  that  from  2  to  3 
per  cent,  of  all  the  sulphur  dioxide  was  converted  into  sulphur  trioxide. 

Lunge''  found  in  two  experiments  with  burning  Spanish  cupriferous 

'  Trans.  A.  I.  M.  E.,  1903,  Vol,  XXXTIL 

2  Trans.  A.  I.  M.  E.,  1905. 

'  "Sulphuric  Acid  and  Alkali  Manufacture." 


pyrites,  containing  48.62  per  cent,  sulphur,  in  a  glass  tube,  in  a  current 

of  air : 

1  2 

Sulphur  obtained  as  SO2,             88 .  02  per  cent.  88 .  78  per  cent. 

Sulphur  obtained  as  SO3,               5 .  80  per  cent.  6 .  05  per  cent. 

Sulphur  in  residue,                         3. 42 per  cent.  1    ^  I7  per  cent. 

Sulphur  lost,                                    2 .  75  per  cent.  / 

Of  the  sulphur  of  the  burner  gas  there  were  present: 

1  2 

As  SO2,  93 .  83  per  cent.         93 .  63  per  cent. 

As  SO3,  6 . 1 7  per  cent.  6 .  37  per  cent. 

Two  experiments  were  made  in  this  way;  in  the  glass  tube  50  grm.  of 
cinders,  from  the  same  pyrites,  in  pieces  of  the  size  of  a  pea,  were  com- 
pletely freed  from  sulphur  by  ignition,  and  fresh  pyrites  burned  as  before, 
the  gas  passing  through  the  cinders.     There  were  found: 

3  4 

Sulphur  as  SO^,  79 .  25  per  cent.         76 .  90  per  cent. 

Sulphur  as  SO3,  16 .  02  per  cent.         16 .  84  per  cent. 

Sulphur  in  residue  and  loss,  4 .  73  per  cent.  6 .  26  per  cent. 

Of  the  sulphur  in  the  burner  gas  itself  there  were  present : 

3  4 

As  SO2,  83 .  18  per  cent.         82 .  00  per  cent. 

As  SO3,  16. 82  per  cent.         18.00  per  cent. 

Experiments  made  by  Scheurer-Kestner  with  gases  from  fur- 
naces roasting  pyrites  for  the  manufacture  of  sulphuric  acid  show  that  the 
sulphur  trioxide  is  quite  variable.  One  set  of  determinations  were 
made  from  two  samples  taken  at  various  times  from  a  lump  kiln  burner, 
and  the  other  from  a  Maletra  fine  ore  burner. 
The  average  results  of  these  experiments  were : 

Vol.  Sulphur  converted 

per  cent.  into  SO3;  per  cent, 

of  SO2  of  total  sulphur 

Lump  burner,  7 . 5  per  cent.  3 . 1  per  cent. 

Fine  ore  burner,  8 . 9  per  cent.  3 . 5  per  cent. 

The  presence  of  sulphuric  acid  in  the  sulphur  fumes,  especially 
those  from  muffle  furnaces,  is  interesting  as  showing  the  formation  of 
sulphuric  acid,  probably  mostly  by  catalytic  action.  This  is  clearly 
shown  in  Lunge's  experiments  Nos.  3  and  4,  where  the  sulphur  dioxide 
was  passed  through  a  column  of  roasted  ore,  mostly  ferric  oxide,  and 
the  amount  of  sulphur  trioxide  increased  from  6.17  per  cent,  and  6.37 
per  cent,  as  shown  in  experiments  1  and  2,  to  16.82  per  cent,  and  18.00 
per  cent.,  as  shown  in  experiments  3  and  4.  The  amount  of  sulphuric 
acid  in  the  fumes  is  also  interesting  from  the  fact  that  the  acid  gases 
absorbed  in  water  has  been  used  as  the  solvent  in  leaching  copper  ores, 
mostly,  however,  after  chloridizing  roasting. 


Sulphur  Determinations. — The  following  method  of  making  sulphur 
determinations  is  used  in  Cripple  Creek  mills,  where  from  30  to  40 
analyses  are  frequently  made  daily.  Instead  of  determining  the  soluble 
and  insoluble  sulphur  from  one  weighing,  two  weighings  are  usually 
made;  one  for  the  total  and  one  for  the  soluble  sulphur.  A  complete  set 
of  samples  for  one  shift  consists  of  an  average  sample  of  the  raw  and  roasted 
ore  from  each  furnace.  In  this  way  a  check  is  kept  on  the  work  done 
by  the  different  shifts. 

The  frequency  with  which  sulphur  determinations  are  made  depends 
upon  the  uniformity  or  changeableness  of  the  ore.  When  ore  is  roasted 
for  a  hydrometallurgical  process  it  is  the  most  important  and  one  of  the 
most  delicate  steps  in  the  entire  treatment,  and  any  indication  of  the 
work  done  by  the  different  shifts  is  desirable. 

Total  Sulphur. — Weigh  1.373  grm.  of  the  finely  powdered  ore  into  a 
No.  4  casserole.  Add  10  c.c.  of  a  saturated  solution  of  potassium  chlorate 
in  nitric  acid.  Cover  with  watch  glass  and  boil  to  dryness.  Add  10  c.c. 
(15  c.c.  for  concentrates)  hydrochloric  acid.  Evaporate  down  to  about 
one-half.  Add  100  c.c.  hot  water  and  boil  slightly.  Add  ammonia  until 
a  precipitate  of  ferric  hydrate  forms,  and  then  add  10  c.c.  of  a  saturated 
solution  of  ammonium  carbonate.  The  ammonium  carbonate  is  to  convert 
any  lead  sulphate  to  carbonate  and  thus  render  the  combined  sulphur 
trioxide  soluble  as  ammonium  sulphate.  Heat  to  boiling,  remove  from 
the  heat,  let  settle,  filter,  and  wash  thoroughly  five  or  six  times.  Acidu- 
late the  filtrate  with  hydrochloric  kcid,  and  then  add  5  c.c.  in  excess. 
Boil,  and  while  boiling  add  a  hot  solution  of  barium  chloride  in  slight 
excess.  Boil  a  few  minutes  longer  and  let  settle.  It  is  best,  before 
filtering,  to  let  the  mixture  remain  hot  or  boiling  slightly,  as  long  as  pos- 
sible, which  greatly  facilitates  the  filtering.  Filter  through  a  9  cm.  fil- 
ter, and  wash  at  least  six  times  with  boiling  water.  Ignite,  and  weigh 
the  barium  sulphate. 

Since  1.373  grm.  were  taken,  the  percentage  of  sulphur  in  the  ore  may 
be  read  directly  from  the  scales,  100  milligrm.  of  barium  sulphate  being 
equal  to  1  per  cent,  of  sulphur  in  the  ore.  In  weighing  out  the  1.373  grm. 
of  ore,  instead  of  making  the  weighings  with  the  usual  gram  and  milligram 
weights,  a  lead  button  or  disc  is  carefully  made  so  as  to  weigh  1.373  grm., 
and  this  lead  button  is  then  always  used  as  the  standard  weight  in  making 
sulphur  determinations. 

If  1/2  grm.  of  ore  is  taken,  as  may  be  desirable  with  ores  high  in 
sulphur,  the  weight  of  the  barium  sulphate  must  be  multiplied  by 
0.1373  to  obtain  the  weight  of  the  sulphur.  To  ignite  the  barium  sul- 
phate, the  filter,  with  the  precipitate,  is  placed  in  an  annealing  cup  and 
heated  to  redness  in  the  muffle.  The  ignited  barium  sulphate  should  be 
perfectly  white. 


In  some  of  the  Cripple  Creek  mills  the  step  of  adding  ammonia  and 
ammonium  carbonate  is  omitted. 

Soluble  Sulphur.— Weigh  1.373  grm.  of  the  finely  pulverized  ore 
into  a  No.  3  casserole.  Add  about  1/2  grm.  sodium  carbonate  and 
20  c.c.  of  water.  Boil  5  minutes.  Remove  and  filter;  wash  thoroughly 
four  or  five  times  with  boiling  water.  Add  10  c.c.  hydrochloric  acid 
to  the  filtrate,  boil,  and  while  boiling  add  a  hot  solution  of  barium 
chloride  in  slight  excess.  Boil  a  few  minutes  longer  and  let  settle. 
Filter  and  wash  at  least  six  times  with  boiling  water.  Ignite  by  placing 
the  filter  and  precipitate  in  an  annealing  cup  and  burn  in  the  muffle  till 
white.     Weigh  as  for  the  total  sulphur. 

Insoluble  Sulphur. — The  insoluble  sulphur  is  determined  by  sub- 
tracting the  soluble  sulphur  from  the  total  sulphur,  by  taking  the  two 
weighings  from  the  same  carefully  mixed  sample,  and  making  a  soluble 
determination  on  one,  and  a  total  determination  on  the  other. 

If  it  is  desired  to  make  the  two  determinations  from  one  weighing 
it  is  first  treated  for  the  soluble  sulphur,  and  then  for  the  insoluble,  by 
treating  it  in  the  same  way  for  the  total  sulphur. 

Tellurium. — Tellurium,  the  analogue  of  sulphur,  is  a  common  asso- 
ciate of  copper,  gold,  and  silver  ores.  In  recent  years  it  has  been  found 
that  this  clement  is  associated  with  gold  in  almost  all  of  the  great  mining 
districts  of  the  world,  even  where  not  long  ago  its  presence  was  unsus- 
pected. It  is  very  common  in  Cripple  Creek,  Colorado,  in  Goldfield, 
Nevada,  and  in  the  Kalgoorlie  mines  of  Australia.  It  occurs,  though 
less  conspicuously,  in  the  black  hills  of  South  Dakota,  in  the  Mount 
Morgan  mine  in  Australia,  and  in  the  San  Juan  mines  of  Colorado.  The 
gold  in  many  of  the  richest  mines  in  the  world  is  associated  with  tellu- 
rium and  sulphur;  the  ore  is  then  known  as  a  sulpho-telluride,  although 
the  tellurium  is  rarely,  if  ever,  found  in  gold,  silver,  or  copper  mining, 
unaccompanied  by  sulphur.  Tellurium  also  occurs  quite  universally 
associated  with  copper,  but  in  quantities  so  minute  as  to  be  of  no  special 
metallurgical  importance.  Many  of  the  copper  ores  of  Arizona,  Butte, 
Montana,  and  of  Australia  contain  small  quantities  of  tellurium — rarely 
exceeding  two  or  three  hundredth  per  cent.  The  matte  from  the  Copper 
Queen,  Arizona,  contains  0.00088  per  cent,  tellurium,  while  that  from 
Butte  contains  from  0.001  to  0.01  per  cent.  The  anode  slimes  from 
electrolytic  copper  refining,  at  Butte,  contains  from  2  to  3  per  cent, 
tellurium  and  selenium,  and  in  some  electrolytic  refineries  the  slimes 
contain  as  high  as  5  per  cent,  of  these  elements. 

In  the  roasting  of  telluride  ores,  the  tellurium  of  itself  is  not  of  any 
great  metallurgical  importance.  At  the  most,  the  quantity  of  it  is  usually 
exceedingly  small  as  compared  with  the  sulphur  and  other  constituents. 
Cripple  Creek  is  widely  known  as  a  telluride  camp,  and  yet  mill  samples 
rarely  show  more  than  a  trace  of  tellurium,  and  frequently  not  that 


Kalgoorlie  is  probably  the  richest  tellurium  district  j'et  discovered,  and 
yet  typical  ore  analyses  show  only  from  0.03  to  0.10  per  cent,  of  tellurium. 

Nevertheless,  tellurium  is  often  a  source  of  anxiety  to  the  metallur- 
gist. In  sulphide  ores  the  metals,  principally  gold,  are  usually  fairly 
evenly  distributed  through  the  rock,  and  the  mineral  contained  in  the 
rock;  but  not  so  in  telluride  ores.  By  far  the  greater  gold  values  are 
intimately  associated  with  the  tellurium,  so  that  where  a  speck  of  tellur- 
ium occurs,  there  is  likely  to  be -associated  with  it  an  appreciable  quantity 
of  gold  also.  Tellurium,  unlike  sulphur,  is  not  usually  evenly  distributed 
through  the  rock.  It  is  ordinarily  concentrated  within  small  areas  and 
cleavage  planes.  The  greater  portion  of  the  gold  in  a  ton  of  telluride 
ore  is  frequently  concentrated  in  a  few  rich  places.  It  is  this  character- 
istic of  telluride  ores,  as  in  ores  containing  free  gold  in  particles  of 
appreciable  size,  that  makes  their  treatment  difficult  by  a  chemical 

Tellurium  fuses  at  500°  C.  (930°  F.)  and  volatilizes  at  a  higher  tem- 
perature (from  550  to  575°  C).  When  roasting  in  an  oxidizing  atmos- 
phere, it  burns  with  a  blue  flame  edged  with  green.  Sylvanite  melts 
easily  tinging  the  flame  greenish-blue.  The  tellurium  combines  with 
oxygen  to  form  tellurium  dioxide  (TeOj)  which  corresponds  with  the 
sulphur  dioxide  (SOj)  formed  in  roasting  sulphides. 

The  gold  compounds  of  tellurium,  usually  sylvanite,  petzite,  and 
calaverite,  all  fuse  at  a  low  heat,  forming  at  first  a  globular  mass,  which, 
when  the  tellurium  is  all  volatilized,  leaves  behind  a  speck  of  gold  of 
definite  proportions — frequently  like  a  pin  head.  If  the  telluride  par- 
ticle is  roasted  at  a  low  temperature,  this  speck  of  gold  will  be  very  porous, 
and  present  a  large  surface  for  attack  by  the  chemical  solvent.  If 
roasted  at  a  sudden  very  high  temperature,  it  is  likely  to  be  round  and 
smooth;  in  this  condition  the  solvents  have  no  appreciable  effect  on  it  in 
the  time  ordinarily  given  by  a  chemical  process.  When,  however,  the 
tellurium  is  largely  associated  with  sulphur,  the  sulpho-telluride  particle 
will  not  fuse,  but  the  sulphur  and  tellurium  will  be  driven  off,  as  in  the 
case  of  sulphides,  leaving  behind  the  gold  disseminated  through  the 
ferric  oxide  particle.  If  the  gold,  in  roasting,  issues  from  its  telluride 
combination  in  a  shotted  form,  the  best  recourse,  after  the  chemical 
treatment,  is  concentration.  This  has  proved  quite  effective.  The 
gold,  although  usually  having  a  bright  yellow  appearance,  does  not  amal- 
gamate well.  Amalgamation  has  been  tried  repeatedly,  but  has  not 
proved  the  success  that  was  anticipated;  nevertheless,  there  seems  to  be 
no  logical  reason  why  it  cannot  be  successfully  accomplished. 

The  difference  in  the  roasting  of  a  grain  of  sulphide  and  a  grain  of 
telluride,  both  of  which  contain  the  same  amount  of  gold,  is  likely  to  be 
this;  in  the  sulphide  the  sulphur  is  gradually  expelled  leaving  the  result- 
ing grain  of  ferric  oxide  extremely  porous  and  with  the  gold  scattered 


through  it  in  perhaps  microscopic  particles;  in  the  telluride,  on  the 
contrary,  unless  the  utmost  care  is  taken,  the  grain  is  likely  to  fuse  into  a 
plastic  mass,  from  which  all  the  gold  contained  in  it  will  finally  emerge 
concentrated  into  one  particle.  The  solvent,  as  subsequently  applied 
to  the  roasted  sulphide  particle,  will  extract  a  very  high  percentage  of 
the  gold  in  a  very  short  time;  while  the  gold  resulting  from  the  roasting 
of  the  telluride  particle  would  scarcely  be  affected.  Some  telluride  ores 
are  easily  treated  and  show  but  little  coarse  gold,  but  the  illustration 
given  shows  why  it  is  usual  to  find  coarse  gold  in  the  tailings  of  telluride 
ores,  even  when  the  chemical  treatment  has  been  apparently  satis- 
factory. It  is  highly  probable  also  that  some  of  the  tellurium,  by  partial 
fusion,  may  be  converted  into  a  compound  which  is  insoluble,  and  which 
resists  further  oxidation  at  higher  temperatures.  This  is  frequently  the 
case  in  the  corresponding  sulphur  combinations,  which  are  insoluble,  and 
from  which  it  is  difficult  to  drive  off  more  sulphur  by  increasing  the  heat. 

Tellurium  is  insoluble  in  water  and  in  dilute  sulphuric  and  hydroch- 
loric acids.  It  is  practically  unaffected  by  chlorine,  bromine,  and  pot- 
assium cyanide.  Gold  and  tellurium  probably  form  true  chemicals 
compounds;  if  this  is  so,  it  is  evident  that  the  gold  cannot  be  closely 
extracted  unless,  in  a  measure,  the  tellurium  is  decomposed.  Gold  tel- 
lurides  are  very  compact  and  do  not  permit  of  much  penetration  by  a  gold 
solvent.  It  is  largely  due  to  these  facts  that  roasting  of  tellurides  is 
desirable  to  get  low  tailings  by  a  solvent  process. 

There  is  no  appreciable  loss  of  gold  by  volatilization  in  the  roasting 
of  telluride  ores,  although  at  very  high  temperatures  some  gold  is  doubtless 
volatilized  with  the  tellurium.  Ordinarily,  however,  the  conditions  in 
roasting  are  such  that  the  tellurium  is  volatilized  before  the  temperature 
is  sufficiently  high  to  volatilize  any  of  the  gold  with  it.  On  comparing 
the  assays  of  the  raw  ore  with  the  roasted  ore,  in  mills  treating  sulpho- 
telluride  ores,  no  loss  by  volatilization  is  apparent  with  an  oxidizing 
roast.  Neither  does  the  dust  in  the  dust  chambers  show  a  much  higher 
value  in  gold  than  the  ore  from  which  the  dust  was  obtained.  The  slight 
difference  in  value  can  be  accounted  for  by  the  fact  that  the  gold  val- 
ues in  the  ore  are  largely  confined  to  the  sulphides  and  tellurides,  and 
owing  to  their  friability  a  larger  proportion  of  the  dust  will  result  from 
these  than  from  the  other  constituents.  If  the  gold  were  appreciably 
volatile  with  the  tellurium,  the  fumes  on  cooling  in  the  dust  chamber 
would  condense  and  appear,  to  some  extent,  in  the  furnace  dust,  and 
there  manifest  itself  in  the  assays. 

Iron. — Iron  is  inseparably  associated  with  copper,  gold  and  silver  ores. 
While  it  is  of  great  importance  mineralogically  and  metallurgically,  it 
presents  no  serious  problems.  In  raw  ore  it  is  frequently  troubelsome 
for  wet  methods,  but  all  difficulties  are  practically  eliminated  by  careful 


Ferric  oxide  (Fefi,),  which  should  be  the  ultimate  condition  of  the 
iron  in  all  roasted  ore,  is  insoluble  and  practically  unaffected  by  all  chem- 
ical solvents  of  copper,  gold,  and  silver.  It  is  immaterial  whether  the 
solvents  are  acid  or  alkaline,  or  whether  they  are  dilute  or  somewhat 
concentrated.  If  iron  gives  any  serious  trouble  in  roasted  ore,  it  is 
entirely  due  to  imperfect  roasting. 

Many  raw  ores  contain  considerable  quantities  of  iron  as  soluble 
sulphate.  This  is  very  pronounced  in  some  mines,  and  particularly, 
in  the  zone  of  partial  oxidation.  When  this  occurs  in  appreciable 
quantities,  the  treatment  of  the  raw  ore  by  any  of  the  chemical  processes 
becomes  difficult,  and  frequently  impossible,  since  all  of  the  chemicals 
used  in  solvent  processes  are  quickly  affected  by  it.  Washing  the  ore, 
either  with  water  or  dilute  acid  or  alkaline  solutions,  is  not  always  effective. 
Roasting  effectively  removes  the  difficulty  by  converting  the  ferrous 
sulphate  into  the  insoluble  ferric  oxide. 

When  copper,  gold,  and  silver  ores  contain  simply  iron  in  any  of  its 
various  combinations  with  the  usual  quartz  matrix,  and  without  appre- 
ciable quantities  of  any  other  foreign  matter,  they  can  be  roasted  quickly, 
thoroughly,  and  cheaply,  no  matter  what  the  sulphur  content  may  be, 
and  the  values  easily  recovered  with  a  high  percentage  of  extraction. 

Iron,  associated  with  copper,  gold,  and  silver  ores  usually  occurs  in 
the  form  of: 

The  Oxide,  Hematite  (Fe203),  with  or  without  the  water  of  hydration. 

The  Sulphide,  Pyrite  (FeS^). 

The  Carbonate,  Siderite  (FeCO,) ,  which,  while  common,  is  not  general. 

When  siderite  is  roasted  it  is  decomposed  according  to  the  reaction: 
FeC03=FeO  +  C02 
and  the  molecule  of  FeO  is  subsequently  converted  to  FeaOg  by  taking 
oxygen  from  the  air.  The  roasted  carbonate  may  be  strongly  mag- 
netic. The  temperature  must  be  carefully  regulated,  to  avoid  sintering 
the  ore,  which  because  of  the  fusibility  of  ferrous  oxide  and  silica,  may 
easily  happen.  According  to  Le  Chatelier,  the  decomposition  of  fer- 
rous carbonate  takes  place  at  800°  C.  (1472°  F.). 

The  oxidized  iron,  in  copper,  gold,  and  silver  veins  is  usually  the 
result  of  the  natural  decomposition  of  the  pyrite  by  aqueous  and  atmos- 
pheric agencies.  With  depth  in  mines,  below  the  influence  of  these 
agencies,  pyritic  ore  will  be  encountered,  while  at  the  surface  the  ore  may 
by  perfectly  oxidized. 

The  principle  objects  to  be  gained  in  roasting  thoroughly  oxidixed 
ores  are  dehydration  and  agglomeration,  which  much  facilitates  the  sub- 
sequent chemical  treatment.  Some  of  the  base  elements  are  invariably 
expelled  or  are  put  in  better  condition  to  resist  the  action  of  the  chemicals. 
Unless  the  iron  is  thoroughly  oxidized  to  the  dehyrated  ferric  oxide, 
much  of  it  is  likely  to  go  into  solution  with  an  acid  solvent. 


Most  of  the  iron  in  oxidized  ores  is  in  the  form  of  ferric  hydrate, 
known  mineralogically  as  limonite  (FezOg+Fe^  (OH)e).  Ores  which 
consist  largely  of  other  substances,  such  as  quartz  and  clay,  usually 
have  the  characteristic  yellow  appearance  of  ferric  hydrate.  By  roasting, 
the  water  of  hydration  is  driven  off,  which  converts  the  iron  into  the 
ferric  oxide  (FejOj).  The  color  at  the  same  time  changes  from  yellow 
or  brown  to  the  familiar  red  of  well-roasted  ore.  In  some  ore  the  red  is 
very  intense. 

The  ferric  hydrate  gives  off  part  of  its  water  at  a  teniperature  between 
80  and  100°  C.  (176  and  212°  F.)  and  all  of  it  at  a  red  heat.  Intense 
heat,  in  roasting  oxidized  ores,  is  not  usually  necessary,  since  sulphates 
in  appreciable  quantities  are  ordinarily  absent.  It  is  the  sulphates 
which  usually  require  a  higher  temperature  for  their  decomposition. 
Oxidized  ores  can  be  quickly  and  cheaply  roasted;  about  all  that  is  neces- 
sary is  to  bring  the  ore  to  a  good  red  heat. 

The  magnetic  oxide,  magnetite  (FcjOJ,  sometimes,  though  not 
frequently,  occurs  associated  with  copper,  gold  and  silver  ores.  When 
roasted  at  a  moderate  temperature,  with  an  abundance  of  air,  it  may 
be  converted  into  the  ferric  oxide.  Roasting  with  salt  appears  to  be 
much  more  effective  in  bringing  about  this  change  than  a  simple  oxidizing 
roast.  Both  magnetic  oxide  and  the  ferric  hydrate  dissolve  to  some  extent 
in  acids,  and  in  a  smaller  degree  are  converted  by  chlorine  into  the 

If  iron  is  contained  in  the  raw  ore  as  pyrite  (FeSj),  the  first 
action  of  the  roasting  is  to  expell  one  atom  of  sulphur.  This  should  be 
accomplished  at  a  moderately  low  temperature  and  with  an  abundance 
of  air.  The  temperature  in  the  early  stages  of  the  roasting  should  not 
exceed  a  dull  red.  As  long,  however,  as  the  ore  does  not  show  any  tendency 
to  adhere  and  form  into  small  lumps  there  is  not  much  danger  of  over- 
heating. It  is  well  known  that  one  of  the  atoms  of  sulphur  in  pyrite  is 
quite  tenaciously  combined  with  the  iron,  while  the  other  is  held  only 
by  a  feeble  bond.  In  an  oxidizing  atmosphere  at  a  temperature  of 
about  315°  C.  (600°  F.)  the  molecule  of  pyrite  beings  to  be  decomposed. 

The  sulphur  in  pyrites,  exposed  to  the  direct  action  of  the  highly 
heated  atmosphere  in  the  furnace,  is  converted  at  once  into  dioxide.  In 
the  deeper  portions  of  the  bed,  where  it  is  difficult  for  the  air  to  pene- 
trate, the  sulphur  may  be  first  volatilized,  and  on  appearing  at  the  sur- 
face, also  burns  to  sulphur  dioxide: 

FeS2  +  02=FeS  +  S02 

S+02  =  S02 

Some  of  the  sulphur  dioxide  in  the  presence  of  large  quantities  of 
incandescent  oxides  or   quartz  is   converted,  by  catalytic  action    into 


sulphur  trioxide,  and  the  sulphur  trioxide  combining  with  the  moisture 
of  the  air  and  fuel  is  converted  into  sulphuric  acid. 

The  ferrous  sulphide,  by  combining  with  the  oxygen  of  the  air,  is 
converted  into  ferrous  oxide: 

FeS  +  03=FeO  +  S03. 

The  ferrous  oxide,  by  combining  with  more  oxygen,  may  be  converted 
mto  the  magnetic  oxide,  or  by  combining  with  the  sulphur  trioxide, 
may  be  converted  into  ferrous  sulphate: 

3FeO  +  0=Fe304 

The  magnetic  oxide,  by  proper  heating  or  by  combining  with  sulphur 
trioxide  may  be  converted  into  ferric  oxide : 

2Fe30,  +  0  =  3Fe203 
2Fe30i  +  S03  =  3Fe203  +  S02 

The  ferrous  sulphate,  at  a  red  heat,  is  decomposed  into  sulphur 
dioxide,  ferric  oxide,  and  ferric  sulphate,  which  on  further  heating  is 
ultimately  decomposed  into  ferric  oxide  and  sulphur  trioxide: 

6FeSO,=Fe2(SOj3  +  2Fe,03+3S02 

The  ultimate  result  of  all  the  reactions  is : 

4FeS2  +  1102  =  2Fe203  +  8S02. 

It  is  desirable  to  convert  all  of  the  iron  into  the  ferric  oxide.  If  the 
heat  is  properly  adjusted,  and  the  ore  remains  in  a  highly  oxidizing 
atmosphere  for  a  prolonged  time,  the  final  result  will  be  the  ferric  oxide. 
If  the  heat  has  been  too  high  or  if  there  has  been  insufficient  air  in  the  fur- 
nace, large  quantities  of  magnetic  oxide  will  be  formed  and  remain  un- 
changed. The  presence  of  magnetic  oxide  in  the  roasted  ore  in  considerable 
quantities  indicates  an  inferior  roast.  Whether  it  is  due  to  the  presence 
of  the  magnetic  oxide  itself,  or  the  condition  which  produced  it,  or  both 
is  difficult  to  determine.  It  has  been  said  that  the  injurious  effects  of 
magnetic  oxide  is  due  to  its  inability  to  resist  the  action  of  the  chemicals  as 
well  as  the  ferric  oxide;  however,  in  the  chlorination  and  acid  processes  the 
consumption  of  acid  or  chlorine  is  less  when  there  is  considerable  magnetic 
oxide  in  the  well-roasted  ore,  nevertheless  the  tailings  are  invariably 
high.  The  presence  of  magnetic  oxide  usually  indicates  a  high  tempera- 
ture roast,  and  ore  roasted  at  a  high  temperature  certainly  resists  the 
action  of  chemicals  better  than  when  roasted  at  a  lower  temperature. 
The  unsatisfactory  extraction  of  high  temperature  roasts  is  probably 
due  to  the  formation  of  silicates  with  iron.     If  the  ore  has  an  unusually 


dark  appearace,  high  tailings  may  be  expected.  The  extraction  is  always 
the  best  when  the  roasted  ore  has  the  red  appearance  of  ferric  oxide. 
The  reaction: 

2Fe304  +  0  =  3Fe203 
by  which  the  magnetic  oxide  is  converted  into  the  ferric  oxide,  is  revers- 

3re203  +  heat  =  2Fe304  +  0 

so  that  if  the  heat  is  too  high  (about  1700  to  1800°  F.) ,  one  atom  of  oxygen 
of  the  ferric  oxide  is  driven  off,  and  the  ferric  oxide  is  converted  into  the 
magnetic  oxide.  This  is  more  likely  to  happen  if  the  atmosphere  in  the 
furnace  is  not  highly  oxidizing,  or  if  the  bed  of  ore  is  too  deep  for  the  air 
to  penetrate,  or  if  the  ore  is  insufficiently  rabbled. 

Magnetic  oxide  is  probably  formed  in  considerable  quantities  in  the 
earlier  stages  of  roasting.  The  color  of  many  ores,  especially  those 
having  considerable  sulphur,  is  quite  dark  while  the  greater  portion  of 
the  sulphur  is  being  eliminated.  It  is  well  known  that  sulphur  dioxide 
is  a  highly  reducing  agent.  The  heated  top  layer  of  the  incandescent 
ore,  as  it  is  turned  over  by  the  rabble,  is  ploughed  under,  so  that  the 
ferric  oxide  particles  are  surrounded  by  the  highly  heated  reducing  atmos- 
phere of  sulphur  dioxide,  which  results  in  reducing  the  ferric  to  the 
magnetic  oxide.     The  reaction  is  probably  represented  by  the  equations: 

FeS,  +  0,=FeS+SO, 

FeS  +  10Fe2O3  =  7Fe3O4  +  SO 

so  that  this  action,  and  the  corresponding  reversible  reaction  by  which 
the  magnetic  oxide  is  reconverted  back  to  the  ferric  oxide,  is  likely  to 
continue  until  the  sulphur  is  pretty  well  eliminated. 

The  greater  part  of  the  iron  sulphide  becomes  sulphate  for  only  a 
very  brief  period,  especially  in  the  early  stages  of  the  roasting.  It  is 
probable  also  that  sulphate  is  not  formed  to  any  very  great  extent  in  the 
ordinary  roast.  Sulphur  determinations  taken  every  hour  in  roasting 
sulphide  ores  show  very  little  soluble  sulphur  in  the  early  stages  of  the 
roasting,  and  at  any  time  only  small  quantities  of  ferrous  sulphate. 
Ferrous  sulphate  is  decomposed  at  about  950°  F.  (510°  C.) 

Ferrous  sulphate  in  roasted  ore  is  highly  injurious,  and  shows  a  very 
defective  roast.  It  is  not  difficult  to  eliminate,  but  it  does  not  follow 
that  when  the  ferous  sulphate  has  all  been  decomposed  that  the  ore  is 
sufficiently  roasted. 

There  are  two  substances  which  offer  simple  chemical  tests  for  iron 
in  solution — potassium  ferricyanide  (FeK3CeN8)  and  potassium  thiocya- 
nate  (KCNS) .  The  ferricyanide  gives  with /errows  salts  a  blue  precipitate 
which  imparts  a  blue  color  to  the  solution,  but  with  ferric  salts  shows  no 
reaction,  but  only  a  brown  color. 


The  test  may  be  made  by  filtering  a  sample  of  the  ore  with 
water,  then  taking  a  drop  of  the  liquid  on  paper  or  on  a  porcelain  plate, 
and  adding  a  drop  or  two  of  the  ferricyanide.  A  blue  color  indicates  the 
presence  of  ferrous  iron.  If  the  iron  is  likely  to  be  in  the  ferric  condition,  as 
in  the  solutions  issuing  from  the  ore  by  the  acid  or  chlorination  processes, 
it  should  be  reduced  from  the  ferric  to  the  ferrous  condition  before 
applying  the  test.  Ferric  salts  are  easily  reduced  to  ferrous  salts  by 
applying  such  reducing  agents  as  zinc,  stanous  chloride,  sulphur  dioxide, 
or  hydrogen  sulphide. 

The  theocyanate  does  not  give  any  marked  coloration  with  ferrous 
iron,  but  with  ferric  iron  in  the  most  dilute  state  it  forms  a  bright  red 
soluble  compound.  The  test  is  made  as  with  the  ferricyanide.  If  the 
iron  is  likely  to  be  in  the  ferrous  condition,  it  should  be  tested  with  the 
ferricyanide,  or  the  ferrous  salt  converted  to  the  ferric  salt  before  apply- 
ing the  test  with  thiocyanate. 

If  no  color  appears,  either  with  the  ferricyanide  or  thiocyanate,  it 
indicates  a  thorough  roast  only  in  so  far  as  the  soluble  sulphur  com- 
pounds of  iron  are  concerned,  but  not  as  to  the  other  constituents  of  the 
ore.  As  a  final  indication  of  the  roast,  except  perhaps  in  pure  pyritic 
concentrates,  or  iron  accompanied  only  by  silica,  these  tests  are  worthless. 

Leaching,  or  boiling,  a  little  of  the  roasted  ore  with  water  and  pre- 
cipitating with  ammonia  will  usually  indicate  the  soluble  iron. 

If  the  ore  is  so  poorly  roasted  as  to  show  undecomposed  sulphides, 
the  roast  is  worthless.  The  best  way  to  ascertain  if  there  is  any  unde- 
composed sulphides  in  the  roasted  ore  is  to  pan  it. 

Copper. — The  mineralogical  combinations  of  copper  are  quite  varied, 
frequently  in  the  same  mine.  It  may  occur  as  the  oxide,  carbonate, 
silicate,  or  sulphide.     The  sulphide  is  by  far  the  most  common. 

In  a  typical  copper  mine,  the  limonite  gossan,  usually  stained  more 
or  less  with  copper,  appears  at  the  surface;  below  the  gossan,  in  the  oxi- 
dized zone,  are  the  oxides,  carbonates,  and  silicates;  then  comes  the 
zone  of  secondary  sulphides  consisting  of  chalcocite,  bornite,  and  chalco- 
pyrite;  and  below  this  the  primary  zone,  consisting  largely  of  pyrite 
interspersed  with  chalcopyrite. 

As  the  oxide,  copper  occurs  as: 
Cuprite,  CujO 
Tenorite,  CuO. 

As  the  carbonate,  it  usually  occurs  as: 
Malachite,  2CuO,  CO^,  H^O 
Azurite,  2CuO,  200^,  H^O. 

As  the  silicate,  it  usually  occurs  as: 
Chrysocolla,  CuSiOj,  2H2O. 


As  the  sulphide,  it  usually  occurs  as: 
Chalcocite,  CujS 
Chalcopyrite,  CuFeSj 
Bornite,  CujFeSg 
Enargite,  CugAsSi 

in  which  the  copper  occurs  as  CujS,  associated  with  antimony  and  arsenic 
sulphides,  and  frequently  with  iron,  lead,  zinc,  and  silver  sulphides. 

The  carbonates,  malachite  and  azurite,  when  roasted  at  a  low  heat 
are  converted  into  cupric  oxide  (black  oxide),  while  the  carbon  dioxide 
and  water  of  hydration  are  driven  off: 

CuCOj,  Cu (OH)  2  +  heat  =  2CuO  +  CO^  +  H^O. 

The  silicate  is  also  converted  at  a  low  red  temperature  to  the  oxide, 
the  color  changing  from  the  characteristic  greenish-blue  of  the  silicate 
to  black. 

Of  the  sulphides  of  copper  only  the  cuprous  sulphide  (CujS)  is  of  any 
metallurgical  importance.  Cupric  sulphide  (CuS)  is  not  stable  at  high 
temperatures,  but  is  decomposed  on  heating  into  cuprous  sulphide  and 
sulphur  dioxide.  When  cuprous  sulphide  is  roasted,  the  copper  is  first 
converted  into  cuprous  oxide  and  sulphur  dioxide: 

Cu,S  +  03  =  Cu20+S02 

and  by  contact  action  is  further  oxidized  into  cupric  oxide  and  sulphur 
trioxide : 

Cu20  +  S02  +  02  =  2CuO  +  S03. 
Some  of  the  sulphur  trioxide,  combining  with  cupric  oxide,  forms  cupric 

CuO+S03  =  CuSO,. 
At  a  higher  temperature,  653°  C,  the  cupric  sulphate  undergoes  decom- 
position, sulphur  trioxide  being  more  or  less  expelled,  so  that  ultimately 
the  sulphate  will  be  reconverted  into  the  oxide. 

If  chalcopyrite,  with  a  quartz  matrix,  is  roasted  at  a  low  heat,  the 
following  reactions  take  place: 

3CuFeS2-l-Heat-hOi8  =  Cu2S+3reSO,  +  CuSO,  +  S02. 
At  590°  C.  the  ferrous  sulphate  decomposes,  and  acting  upon  the 
cuprous  sulphide  remaining,  converts  it  into  the  cupric  sulphate: 

Cu,S+FeS0,  +  0e  =  2CuS0,+FeA- 
At  650°  C.  the  copper  sulphate  is  decomposed  into  basic  sulphate 
and  sulphur  trioxide,  and  at  700°  C,  into  cupric  oxide  and  sulphur 
trioxide  as  follows: 

2CuSO,=  CuO,CuSO,-i-S03 
CuO,  CuSO,  =2CuO  +  S03. 


The  ultimate  condition  of  the  roasted  product,  therefore,  when  carried 
to  above  700°  C.  are  ferric  oxide,  and  cupric  oxide.  If  the  temperature 
is  not  carried  above  700°,  sulphate  of  copper  may  remain,  while  if  the 
temperature  is  not  carried  above  the  decomposition  point  of  ferric 
sulphate,  both  copper  and  iron  sulphates  will  remain  in  the  roasted  ore. 
For  a  sulphatizing  roast,  the  temperature  should  not  exceed  650°  C. 

Cuprous  sulphide  fuses  readily,  and  if  contained  in  the  ore  in  any 
considerable  quantity,  must  be  heated  carefully  to  avoid  fusing. 

As  long  as  sulphur  dioxide  is  being  produced  by  the  oxidation  of  the 
sulphur,  cupric  oxide  (CuO)  cannot  be  formed.  As  soon  as  all  the  sul- 
phide is  converted  into  a  mixture  of  cuprous  oxide  and  sulphate,  the 
cuprous  oxide  begins  to  be  converted  into  the  cupric  oxide;  and  if  the 
roasting  is  continued  long  enough,  all  the  copper  in  the  ore  will  be  con- 
verted into  the  cupric  oxide,  with  the  probable  formation  of  silicates 
also,  if  there  is  silica  present. 

With  a  low  temperature,  the  copper  may  be  contained  in  the  roasted 
ore  as  cupric  oxide  (CuO),  cuprous  oxide  (CU2O),  and  cupric  sulphate 
(CuSOJ,  and  this  is  usually  the  best  condition  for  copper  extraction, 
provided  there  are  no  deleterious  sulphates  in  the  roasted  ore,  or  undecom- 
posed  sulphides.  If  the  ore  is  poorly  roasted,  some  of  the  sulphides  may 
remain  undecomposed,  and  this  would  unfit  it  for  a  solvent  process.  If 
undecomposed  sulphides  are  suspected,  it  is  best  to  pan  some  of  the 
roasted  material,  when  the  sulphides  will  be  made  apparent. 

If  the  amount  of  copper  in  the  ore  is  small,  as  for  example  in  cup- 
riferous gold  and  silver  ores,  all  the  copper  sulphide  will  be  converted 
into  the  cupric  oxide  when  roasting  sufficiently  low  in  sulphur  to  make 
the  ore  suitable  for  a  solvent  process,  provided  the  finishing  heat  has 
not  been  excessive.  Whatever  the  condition  of  the  copper  in  the  raw 
ore,  in  the  roasted  ore  it  will  appear  as  cupric  oxide  if  the  ore  has  been 
properly  roasted. 

If  the  ore  is  to  be  treated  principally  for  its  copper  content,  it  will 
not  ordinarily  be  necessary  nor  desirable  to  roast  to  such  a  complete 
state  of  oxidation  as  in  ore  treated  principally  for  the  precious  metals. 
For  the  acid  processes,  sulphate  of  copper  is  riot  harmful,  and  is  usually 
highly  beneficial,  but  no  sulphides  should  be  in  evidence.  If  the  same 
ore  is  to  be  treated  for  gold  or  silver  by  the  chlorination  process  the 
preliminary  acid  treatment  for  the  extraction  of  the  copper  puts  it  in 
the  best  possible  condition  for  the  extraction  of  the  gold  and  silver.  If 
cyanidation  is  to  follow  the  acid  treatment,  a  thorough  alkaline  wash  is 

The  greatest  danger  in  roasting  copper  ores  is  in  the  early  stages  of 
the  process.  Cuprous  sulphide,  as  previously  stated,  is  readily  fusible. 
Cuprous  oxide  melts  at  a  red  heat,  but  the  cupric  oxide  is  quite  infusible. 
Cupric  oxide  is  more  readily  soluble  than  cuprous  oxide.     If  the  tempera- 


ture  is  excessive  during  the  roasting,  ferrites  (CuOjFejOj)  and  silicates 
are  likely  to  form,  and  the  copper  in  these  combinations  is  soluble  only 
with  the  greatest  difficulty.  It  will  usually  be  more  satisfactory  to 
slightly  under-roast  copper  ore  than  to  take  chances  in  getting  the  best 
possible  roast  by  overheating.  Copper  ores  are  particularly  sensitive 
to  high  temperatures,  that  is  to  say,  temperatures  above  a  dull  red  (650° 
C.  or  1202°  F.)  and  if  sintering  or  fusion  occurs  it  is  practically  impos- 
sible to  get  a  satisfactory  extraction.  Cupric  oxide  is  reduced  to  cu- 
prous oxide  at  1050°  C. 

The  best  way  to  determine  the  best  conditions  of  roasting,  prin- 
cipally as  to  the  temperature,  is  by  direct  experimenting,  and  leaching 
the  roasted  ore  with  dilute  hydrochloric  or  sulphuric  acid.  If  the 
roasting  has  been  properly  done,  there  should  be  no  difficulty  in  extract- 
ing at  least  90  per  cent,  of  the  copper.  If,  however,  the  temperature 
has  been  excessive,  a  very  poor  extraction  of  the  copper  may  be  expected. 
In  making  preliminary  tests,  it  is  well  to  roast  at  least  one  lot  of  ore  at  a 
temperature  no  higher  than  scarcely  a  visible  red,  and  then  increase  the 
temperature  on  successive  charges.  The  precentage  of  copper  extracted 
in  a  certain  reasonable  time  from  the  different  roasted  samples  will, 
by  comparison,  give  the  highest  temperature  the  ore  will  stand  without 
detriment,  and  that  is  the  temperature  at  which  the  ore  should  be 

Whether  or  not  there  is  any  copper  sulphate  in  the  roasted  ore  can 
easily  be  ascertained  by  placing  a  small  portion  in  a  funnel,  leaching  it 
with  hot  water,  and  then  adding  ammonia  to  the  filtrate,  when,  if  there 
is  any  soluble  copper,  the  familiar  blue  will  appear.  If  no  blue  appears, 
it  may  be  assumed  that  the  copper  in  the  roasted  ore  is  all  in  the  condition 
of  oxide. 

In  gold  and  silver  ores,  where  the  copper  does  not  occur  in  sufficient 
quantities  to  attempt  to  recover  it  at  a  profit,  the  principal  injurious 
effect  of  the  copper  is  in  the  consumption  of  chemicals,  and  in  its  pre- 
cipitation with  the  precious  metals.  It  is  also  undesirable  in  the  gold 
and  silver  bullion,  unless  there  is  sufficient  copper  to  make  electrolytic 
refining  possible.  The  extent  to  which  copper  in  the  ores  of  the  precious 
metals  is  fatal  will  depend  largely  on  the  price  of  the  chemicals,  on  the 
consumption  of  chemicals,  and,  in  a  measure,  on  the  roast.  The  best 
ultimate  condition  of  the  copper  in  roasted  ore  of  the  precious  metals  is 
in  the  form  of  oxide,  and,  fortunately,  this  is  the  way  it  usually  occurs. 

If  the  ore  is  roasted  with  salt,  much  of  the  copper  will  be  converted 
into  cupric  chloride  (CuClj),  but  since  at  a  red  heat  cupric  chloride  gives 
up  half  its  chlorine,  some  of  the  copper  will  be  in  the  form  of  cuprous 
chloride  (CuCl) .  If  the  temperature  is  very  high  most  of  .the  cupric 
chloride  will  be  converted  into  the  cuprous  chloride  and  oxychloride 
of  copper,  both  of  which  are  wholly  or  partly  insoluble  in  water.     The 


chloridizing  roasting  of  copper  ores  is  taken  up  more  fully  under  "Chlo- 
ridizing  Roasting." — "Longmaid-Henderson  Process." 

The  ideal  roast  for  copper  ores  is  one  in  which  all  the  copper  is  in  the 
form  of  sulphate  and  oxide,  and  all  the  iron  in  the  ferric  condition.  This 
represents  a  roast  in  which  practically  all  the  copper  is  soluble  in  water  or 
dilute  acids,  and  all  the  iron  insoluble.  Since  ferrous  sulphate  is  decom- 
posed at  590°  C,  and  curpic  sulphate  at  650°  C,  the  best  roast  for  the  cop- 
per, theoretically  at  least,  should  be  obtained  by  maintaining  the  tempera- 
ture between  590  and  650°  C.  This  also  represents  the  best  conditions 
in  practice. 

The  temperature  at  which  ferrites  and  silicates  of  copper  are  formed 
has  not  been  definitely  determined,  but  if  the  ore  is  heated  much  above 
700°  C,  there  is  great  danger,  and  the  ferrites  and  silicates  once  formed, 
the  satisfactory  extraction  of  the  copper  presents  a  problem  of  some 
magnitude.  The  presence  of  ferrites  and  silicates  is  usually  indicated  by 
the  dark  appearance  of  the  ore,  instead  of  the  red  color  of  ferric  oxide. 

The  sulphur  elimination  in  roasting  copper  ores  depends  largely 
on  the  amount  of  copper  in  the  ore,  as  well  as  on  the  amount  of  sulphur. 
If  much  of  the  copper  in  the  roasted  ore  is  soluble,  it  is  soluble  as  the 
sulphate,  and  the  soluble  sulphur  will  be  quite  large.  In  this  respect 
roasting  of  copper  ores  differs  somewhat  from  roasting  gold  and  silver 
ores,  where  the  sulphur  content  of  the  roasted  ore  is  necessarily  quite 
low.  It  is  not  desirable  to  eliminate  the  same  amount  of  sulphur  from 
copper  ores  as  from  gold  and  silver  ores,  so  that  the  roasting  of  copper 
ores  suitable  for  hydrometallurgical  extraction  will  usually  be  somewhat 
cheaper  than  if  the  same  ore  were  roasted  for  the  extraction  of  the 
precious  metals.  On  the  other  hand,  copper  ores  usually  require  a 
more  delicate  roast  than  gold  and  silver  ores. 

For  high  grade  cupriferous  concentrates  or  sulphide  ore,  the  roasted 
product  may  contain  from  3  to  5  per  cent,  sulphur,  and  be  well  roasted 
for  chemical  treatment. 

It  is  possible  to  make  most  of  the  copper  soluble  by  prolonged  roasting 
at  a  low  temperature  and  with  insufficient  air,  but  such  a  roast  will  also 
leave  much  of  the  iron  and  other  constituents  soluble.  Whether  this 
is  desirable  or  not  depends  almost  entirely  on  the  method  of  precipitating 
the  copper,  and  on  the  chemical  composition  of  the  soluble  iron. 

In  cases  where  gold  and  silver  ores  contain  copper,  the  roasting 
must  be  effected  to  get  the  best  extraction  of  those  metals,  but  if  a  little 
care  is  used  in  the  temperature  of  the  roasting,  a  high  extraction  of  the 
copper  may  also  be  made,  even  if  the  ore  is  roasted  with  a  view  of  getting 
the  best  extraction  of  the  gold  and  silver. 

It  is  difficult  to  understand  why  some  sulphide  ores  with  careful  roast- 
ing will  give  up  their  values  while  with  others  of  apparently  the  same 
or  similar  composition  it  is  difficult  to  get  a  reasonable  extraction. 


In  order  to  determine  the  effect  of  roasting  on  chalcopyrite,  per  se, 
several  pounds  of  the  pure  mineral,  containing  as  a  matrix  pure  quartz 
and  some  galena,  was  carefully  roasted  in  an  assay  muffle.  The  heat 
was  never  above  a  dull  red.  The  copper  content  of  the  raw  ore  was  29.3 
per  cent.,  and  the  sulphur  28.4  per  cent.  The  sulphur  in  the  roasted  ore 
was  5.5  per  cent.,  of  which  3.5  per  cent.,  was  soluble.  The  high  insoluble 
was  probably  largely  due  to  the  galena,  forming  lead  sulphate. 

The  extraction,  by  agitating  with  a  5  per  cent,  sulution  of  sulphuric 
acid,  was  as  follows: 

Raw  ore,  29 . 3  per  cent.  Cu. 

Roasted  ore,  1  hour's  treatment,  3 . 2  per  cent.  Cu.  Extraction,  89 . 1  per  cent 

2  hours'  treatment,  2.2  per  cent.  Cu.  Extraction,  92.5  per  cent 

4  hours'  treatment,  1 . 2  per  cent.  Cu.  Extraction,  95 . 9  per  cent 

On  the  other  hand,  in  just  as  carefully  roasting  a  sulphide  concen- 
trate from  Mexico,  containing  6  per  cent,  copper,  it  was  difficult  to  get 
an  extraction  of  80  per  cent.  The  mineralogical  combination  of  the 
copper  in  these  concentrates  was  not  determined,  but  presumably  it  was 
in  the  form  of  chalcopyrite  or  bornite. 

Silver. — Silver  is  universally  associated  with  copper  and  gold.  It 
may  be  said  that  gold  in  ores  is  never  found  unaccompanied  by  silver, 
while  copper  is  a  common  associate  of  both  gold  and  silver. 

Silver  is  an  important  factor  in  the  treatment  of  copper  and  gold 
ores.  The  silver  is  not  readily  recovered,  and  if  contained  in  the  tailings 
in  any  considerable  quantity,  may  act  as  an  obstacle  to  the  close  extrac- 
tion of  the  gold,  the  fact  that  silver  is  not  readily  soluble  in  any  of  the 
ordinary  solvents  of  copper  and  gold  adds  somewhat  to  the  difficulty  of 
its  extraction  by  the  wet  processes.  Roasting,  in  the  metallurgy  of 
silver,  is  a  very  important  factor. 

If  there  is  any  free  gold  in  the  ore,  the  silver  will  be  more  or  less 
alloyed  with  the  gold.  In  thoroughly  oxidized  ores  it  probably  occurs 
in  this  way.  In  unoxidized  ores  it  will  almost  always  be  found  as  the 
sulphide,  associated  more  or  less  with  arsenic,  antimony,  and  copper,  as 
will  be  noticed  from  the  common  minerals  of  silver: 

Argentite  (silver  glance),  AgjS. 

Pyrargyrite  (ruby  silver),  3Ag2S,Sb2S3 

Proustite,  2Ag2S,As2S3 

Stephanite,  5Ag2S,Sb2S3 

Stromeyerite,  Ag2S,Cu2S 

Polybasite,  _  9(Ag2S,Cu2S),Sb2S3,As2S3 

Cerargyrite  (horn  silver),  AgCl 

Hessite,  Ag2Te 

Petzite,  (Ag,Au)2Te. 


While  silver  is  more  or  less  associated  with  tellurium,  it  is  a  strange 
fact  that  only  a  comparatively  small  quantity  of  silver  is  found  associated 
with  gold  in  telluride  or  sulpho-telluride  ores,  and  is  not  of  any  great 
consequence  either  in  the  oxidizing  roasting  or  subsequent  chemical 
treatment.  When  it  does  occur  combined  with  gold  and  tellurium,  the 
tellurium  is  volatilized  in  roasting,  leaving  behind  an  alloy  of  gold  and 

In  the  oxidizing  roasting  of  sulphide  copper  and  gold  ores  containing 
silver,  the  silver  sijlphide  is  first  converted  into  the  sulphate  at  an  early 
stage  of  the  operation.  The  silver  sulphide  reacting  with  the  sulphur 
trioxide,  formed  principally  from  iron  and  copper  sulphides  by  catalytic 
action  with  hot  silica  and  metal  oxides,  forms  silver  sulphate,  and  reduces 
the  sulphur  trioxide  to  the  dioxide: 

2Ag2S  +  4SO3  =  Ag,SO,  +  4S0,. 

If  there  is  any  free  gold  in  the  sulphide  ore,  as  there  frequently  is  in 
small  quantities,  some  silver  will  also  be  free  but  alloyed  with  the  gold. 
This  silver  in  the  first  stages  of  roasting  is  likely  to  be  converted  into  the 
sulphate : 

2Ag  +  2S03  =  Ag,SO,  +  S02. 

As  the  ore  is  rabbled  against  a  higher  temperature  and  a  more  highly 
oxidizing  atmosphere,  the  silver  sulphate  is  gradually  converted  into 
metallic  silver.  The  silver  sulphate  is  partly  reduced  by  the  direct 
action  of  heat  alone: 

Ag2S0,+  Heat  =2Ag  +  S03  +  0 

but  the  temperature  required  for  this  reaction — 1095°  C. — is  rarely  if 
ever  attained  in  a  roasting  furnace.  In  the  presence  of  reducing  gases, 
silver  sulphate  is  decomposed  at  a  very  moderate  heat,  metallic  silver 
being  deposited.  In  the  presence  of  copper  oxides,  silica,  and  iron  oxides, 
silver  sulphate  is  decomposed  at  temperatures  from  860  to  870°  C. : 

Ag2SO,  +  4Fe304  =  2Ag  +  6Fe203+S02 
Ag2S04  +  Cu20    =2Ag  +  CuS04  +  CuO. 

In  the  roasting  of  copper,  gold,  and  silver  ores  suitable  for  treatment 
by  chemical  processes,  in  a  highly  oxidizing  atmosphere,  the  silver  in  the 
ore  will  always  be  found  in  the  metallic  condition,  alloyed  with  the  gold. 
It  is  for  this  reason  that  silver  plays  so  important  a  part  in  the  hydro- 
metallurgical  treatment,  if  alloyed  in  appreciable  quantities  with  gold. 
Unless  the  extraction  of  the  silver  is  quite  thorough  there  may  enough 
remain  in  the  tailings  to  protect,  ia  a  measure,  the  gold  from  the  action 
of  the  solvent. 

If  the  sulphide  ore  is  improperly  roasted,  some  of  the  silver  may 


remain  as  sulphate.     As  such  it  is  readily  soluble  in  water.     One  hundred 
parts  of  water  dissolves  0.58  parts  of  silver  sulphate. 

It  has  been  conclusively  proved  that  silver  is  volatilized  in  oxidizing 
roasting.  It  is  possible  that  in  many  instances  the  volatilization  of 
silver  in  oxidizing  roasting  is  due  to  the  presence  of  small  quantities  of 
cerargyrite,  or  natural  silver  chloride.  The  chloride  of  silver  seems  to  be 
quite  generally  distributed  in  the  various  silver  ores.  The  chloride 
volatilizes  at  a  strong  red  heat,  so  that  if  an  excessive  loss  of  silver  is 
discovered  in  oxidizing  roasting  it  is  well  to  examine  the  ore  to  ascertain 
the  presence  of  chlorine. 

Since  metallic  silver  is  with  difficulty  soluble  by  any  of  the  commercial 
processes  for  the  recovery  of  copper,  gold  and  silver,  chloridizing  roasting 
is  frequently  resorted  to,  in  order  to  convert  the  silver  into  the  more 
soluble  silver  chloride.  This  subject  is  taken  up  in  detail  under  the  head 
of  "Chloridizing  Roasting." 

Any  chemical  process  having  for  its  primary  object  the  recovery  of 
copper  or  gold  from  its  ores  must  take  cognizance  of  the  silver  usually 
associated  with  them.  If  the  quantity  of  silver  is  small  there  is  no 
difficulty  in  recovering  a  fair  percentage  by  either  the  cyanide  or  chlorina- 
tion  processes.  If  the  quantity  is  large,  the  best  average  results  will  be 
obtained  by  chloridizing  roasting,  when  the  silver  may  be  extracted  quite 
closely  by  either  the  hyposulphite,  cyanide,  or  chlorination  processes. 

Gold. — Gold,  of  itself,  is  of  no  metallurgical  importance  in  the  process 
of  oxidizing  roasting.  It  always  occurs  native  or  mixed  or  combined 
with  sulphur  or  tellurium;  but  whether  mixed  or  combined,  on  roasting  it 
emerges  as  metallic  gold,  which  at  all  stages  of  the  roasting  is  unaf- 
fected by  any  temperature  or  condition  of  the  furnace.  If  the  gold  is 
free  and  microscopically  fine,  coarse,  flaky,  solid,  or  porous,  it  will,  of 
itself,  remain  so.  It  will  appear  in  the  finally  roasted  ore  as  it  appeared 
in  the  raw  ore,  or  after  being  liberated  in  the  early  stages  of  the 

Much  has  been  said  about  the  loss  of  gold  in  roasting.  Careful  inves- 
tigation has  shown  that  in  oxidizing  roasting  thei'e  is  none  but  a  mechan- 
ical loss,  which  is  subject  to  the  same  conditions  as  the  handling  of  dry 
ore  under  any  circumstances.  If  the  ore  contains  silver  chloride,  which 
usually  has  associated  with  it  some  gold,  a  loss  of  gold  may  be  expected 
with  that  of  the  silver,  in  oxidizing  roasting,  but  this  loss  will  ordinarily 
be  very  small  if  the  temperature  is  properly  regulated,  and  in  no  case 
will  it  ever  be  serious. 

Lead. — Lead  usually  occurs  as  the 

Sulphide  (PbS),  Galena, 
Carbonate  (PbCOg),  Cerussite, 
Sulphate  (PbSOJ,  Anglesite. 


Lead  is  more  or  less  associated  with  copper  and  gold  ores,  but  not 
usually  in  large  quantities.  Silver  is  more  commonly  associated  with 
it.  The  presence  of  lead  in  ores  in  small  amounts  is  not  particularly 
harmful,  either  in  roasting  or  in  the  subsequent  treatment  by  the  hydro- 
metallurgical  processes.  When  lead  occurs  in  ores  in  large  quantities,  it 
is  so  desirable  as  a  smelting  material  that  its  treatment  by  solvent  process 
is  quite  remote. 

According  to  Plattner,  if  the  sulphide,  galena,  is  roasted  at  a  low 
temperature,  to  prevent  fusion,  it  will  at  first  be  converted  into  the 
oxide  and  sulphur  dioxide: 

PbS  +  03=PbO+S02. 

A  part  of  the  sulphur  dioxide  on  coming  in  contact  with  the  heated 
silica  combines  with  the  oxygen  of  the  air  to  form  sulphur  trioxide,  and 
this  combines  with  the  lead  oxide  to  form  lead  sulphate : 

PbO  +  S03=PbSO,. 

The  lead  in  the  roasted  ore  will  usually  be  in  the  form  of  oxide  and 
sulphate.  Silicate  will  not  occur  unless  the  ore  has  been  fused,  in  which 
case,  in  any  event,  the  ore  would  probably  be  unfit  for  subsequent 
chemical  treatment.  The  proportion  of  the  oxide  to  the  sulphate  will 
depend  upon  the  presence  of  other  sulphides,  the  method  of  roasting, 
and  to  a  large  extent  on  the  proportion  of  the  galena  to  the  other  con- 
stituents in  the  ore. 

It  is  probable  that  much  of  the  sulphide  may  be  converted  directly 
into  the  oxide  and  sulphate  by  the  slow  roasting  and  at  the  low  temper- 
ature usually  employed  in  roasting  copper,  silver,  and  gold  ores  for 
subsequent  chemical  treatment: 

2PbS  +  O7  =  PbSO  4 + PbO  -h  SO  2. 

"When  the  ore  contains  considerable  lime,  as  sometimes  happens, 
some  metallic  lead  may  possibly  be  formed: 

4PbS+4CaO=3CaS  +  CaS04-t-4Pb. 
CaS-I-0,  =  CaSO,. 

The  transformation  of  the  calcium  oxide  to  the  calcium  sulphate 
is  a  desirable  change  for  the  subsequent  chemical  treatment.  Quartz, 
clay,  and  silicates  remain  inert  to  lead  sulphide. 

Galena  is  difficult  to  roast.  It  fuses  at  a  low  temperature,  and  if 
excessively  heated,  is  likely  to  agglomerate  into  a  mixture  of  lead  oxide 
and  lead  sulphate  from  which  it  is  difficult  to  expel  the  sulphur  trioxide, 
even  when  heated  so  high  as  to  melt,  and  from  which  it  is  absolutely 
impossible  to  satisfactorily  extract  the  copper  and  precious  metals. 
Lead  sulphate,  with  heat  alone,  is  decomposed  only  at  a  white  heat. 


Lead  carbonate  (PbCOa)  is  readily  decomposed  at  a  low  temperature 
(200°  C;  392°  F.)  into  lead  oxide  and  carbon  dioxide: 


By  prolonged  roasting,  lead  carbonate,  or  the  monoxide,  at  a  temper- 
ature of  not  exceeding  450°  C.  (842°  F.)  may  be  converted  into  the  higher 
oxide,  minium  (PbaOJ.  At  a  still  higher  temperature  the  red  lead 
or  minium  again  gives  up  its  oxygen  and  is  reconverted  into  the  monoxide 
or  litharge. 

The  lead  in  roasted  ore,  on  cooling,  is  likely  to  be  in  the  form  of  mon- 
oxide (PbO);  rarely  perhaps  as  the  red  oxide  (PbgOJ;  some  sulphate 
(PbSOJ,  and,  if  the  ore  is  fused,  as  silicate. 

Ore  containing  as  high  as  10  per  cent,  lead,  can  with  care,  be  satis- 
factorily roasted  for  subsequent  chemical  treatment. 

Zinc. — Zinc  frequently  occurs  associated  with  copper,  gold,  and 
silver  ores  as: 

Sulphide  (ZnS),  Sphalerite. 
Oxide  (ZnO),  Zincite. 
Carbonate  (ZnCOj),  Smithsonite. 
Silicate  (Zn^SiO^),  Willemite. 

In  the  oxidized  ores  the  zinc  usually  occurs  as  the  oxide  or  carbonate; 
in  the  sulphide  ores  it  is  always  found  as  sphalerite. 

If  the  zinc  is  in  the  form  of  carbonate,  roasting  readily  drives  off  the 
carbon  dioxide,  leaving  the  oxide  of  zinc; 

ZnC03  =  ZnO+C02. 

If  the  zinc  is  in  the  form  of  sulphide,  oxidation  at  the  temperature  at 
which  copper,  gold,  and  silver  ores  are  usually  roasted,  takes  place 
slowly,  and  yields  a  mixture  of  oxide  and  sulphate.  The  amount  of 
sulphate,  however,  is  small  as  compared  with  the  oxide. 

Zinc  sulphide  begins  to  oxidize  at  a  dull  red  heat.  As  the  temperature 
is  increased  the  oxidation  takes  place  more  rapidly,  with  the  formation 
of  zinc  oxide  and  sulphur  trioxide: 

ZnS  +  03  =  Zn0  +  S02. 
2ZnS  +  07  =  ZnO  +  ZnS04  +  SOj. 

At  a  prolonged  high  temperature  the  sulphate  is  converted  into  the 

ZnS04  +  heat  =  ZnO  +  S03. 

Some  of  the  sulphur  dioxide  released  in  roasting  by  catalytic  action 
with  the  glowing  ore  is  converted  into  the  trioxide,  which  may  then  com- 
bine with  some  of  the  zinc  oxide  to  form  the  neutral  sulphate,  ZnS04. 



In  the  decomposition  of  the  neutral  sulphate,  basic  sulphates  may  be 
formed  which  require  a  high  and  prolonged  temperature  to  ultimately 
resolve  them  into  the  oxide.  Zinc  sulphate,  by  heat  alone,  is  decomposed 
at  739°  C. 

Zinc  sulphide  does  not  oxidize  as  readily  as  iron  or  copper  sulphides, 
and  a  higher  temperature  is  required  to  start  oxidation.  It  is  infusible 
at  any  temperature  attained  in  the  roasting  furnace,  in  roasting  copper, 
gold  and  silver  ores.  Its  presence,  so  far  as  roasting  is  concerned,  is  not 

The  relative  proportion  of  zinc  oxide  and  zinc  sulphate  formed  in 
roasting  will  depend  upon  the  temperature,  the  oxidizing  qualities  of  the 
atmosphere,  and,  a  certain  extent,  the  relative  quantity  of  zinc  in  the  ore. 
The  higher  the  temperature  in  a  highly  oxidizing  atmosphere,  the  more 
zinc  oxide,  and  the  less  zinc  sulphate,  will  be  formed. 

Zinc,  though  quite  generally  distributed,  is  not  frequently  found  in 
copper,  gold,  and  silver  ores  in  sufficient  quantity  to  interfere  with  the 
metallurgical  treatment  by  wet  methods.  If  contained  only  in  small 
amounts,  and  the  ore  is  thoroughly  roasted,  practically  all  the  zinc  will 
be  in  the  form  of  oxide.  The  oxide  is  readily  soluble  in  acids,  but  is 
not  so  readily  acted  upon  by  chlorine,  cyanide  or  sodium  hyposulphite, 
although  it  affects  these  solvents  injuriously. 

Zinc  oxide  is  not  volatile  at  the  highest  temperature  used  in  roasting 
ores— from  900  to  970°  C.  (1652  to  1778°  F.).  Zinc  in  its  metallic  con- 
dition is  quite  volatile,  even  at  a  moderately  low  temperature.  Reducing 
gases,  such  as  carbon  monoxide  from  the  fuel,  have  a  tendency  to  re- 
duce the  oxide  to  the  metallic  zinc  and  thus  volatilize  it: 

ZnO+CO  =  Zn-t-CO,. 


(H.  O.  Hofman,  Trans.  A.  I.  M.  E.,  1905) 

Heated  ia 



Carbon  dioxide. 
Carbond  dioxide 

deg.  C. 


Total  S. 

1 .  63  per  cent 
1 .  79  per  cent. 
0 .  50  per  cent. 
0 .  50  per  cent. 

S.  eliminated 
as  SO, 

0.14  per  cent. 
0.14  per  cent. 
0.18  per  cent. 
0.16  per  cent. 

S.  eliminated 
as  so. 

1 .  49  per  cent. 
1 .  65  per  cent. 
0.32  per  cent. 
0 .  34  per  cent. 

Ratio . 
S  as  SO2 
S  as  SO  3 



In  these  experiments,  the  temperature  of  the  furnace  was  brought 

slowly  to  the  point  at  which  the  first  acid  was  given  off,  then  raised  about 

10°  C.  and  maintained  constant  for  several  hours.     While  only  a  very 

small  proportion  of  the  total  sulphur  was  driven  off  in  the  tests,  they 



show  that  in  heavy  zinc  sulphate  the  tendency  of  the  salt  to  split  into 
ZnO,  SO2,  and  O  is  greater  when  oxygen  is  absent  than  when  it  has  free 
access  of  air.    . 

Arsenic. — Arsenic  usually  occurs  associated  with  copper,  gold,  and 
silver  ores  as  the  sulphide: 

Arsenopyrite   (FeAsS),  Mispickel. 
'  Realgar  (AsjSj) 

Orpiment  (AsjSg) 

Arsenic  is  almost  universally  associated  with  sulphide  copper  ores. 
If  occurring  in  small  quantities  it  is  of  no  special  importance  in  the 
roasting  except  that  its  elimination  should  be  as  complete  as  possible 
under  the  conditions  of  the  roast.  Its  presence  in  the  precipitated  copper 
is  very  harmful,  and  hence  effort  should  be  made  to  keep  the  solutions  as 
free  from  it  as  possible. 

The  sulphides  of  arsenic  fuse  readily  so  that  care  must  be  exercised 
in  the  first  stages  of  the  roasting.  Arsenic  volatilizes  at  a  comparatively 
low  temperature,  in  the  condition  of  arsenous  oxide.  In  the  presence  of 
an  excess  of  oxygen  there  is  a  tendency  to  form  arsenates  of  iron  and  other 
metals,  and  these  arsenates  are  decomposed  only  at  an  exceedingly  high 
temperature.  After  the  arsenic  is  driven  off  in  the  first  stages  of  the 
roasting,  no  harm  can  result  in  elevating  the  temperature  to  that  re- 
quired for  the  other  constituents  of  the  ore.  Arsenates  are  undesirable 
in  the  roasted  ore,  as  they  interfere  with  the  close  extraction  of  the  copper 
and  precious  metals.  Arsenic  in  the  ore,  as  a  rule,  is  not  particularly 
detrimental  to  any  of  the  solvent  processes,  if  the  roasting  is  properly 

Mr.  R.  R.  Rothwell,  in  speaking  of  roasting  arsenical  pyrites  at 
Deloros,  Canada'  says,  "It  was  asserted  by  some  metallurgists  that  the 
roasting  of  arsenical  pyrites  presents  many  difficulties.  I  can  affirm,  on 
the  contrary,  that  they  roast  with  much  greater  facility  and  in  about 
two-thirds  of  the  time  necessary  to  roast  simple  sulphides.  They  stand 
almost  any  amount  of  heat  without  fusing,  and  the  arsenic,  wljich  forms 
about  49  per  cent,  of  the  mispickel,  volatilizes  at  comparatively  low 
temperature,  seems  to  leave  the  mass  porous,  thus  facilitating  the  oxida- 
tion of  the  sulphur."  An  extraction  of  95  per  cent,  of  the  gold  was  made 
on  these  arsenical  pyrites  by  the  chlorination  process. 

At  Murcur,  Utah,  where  the  ore  has  been  treated  successfully  for 
many  years  by  the  cyanide  process,  roasting  has  been  found  to  eliminate 
any  injurious  effects  from  the  arsenic,  which  occurs  in  considerable 
quantities  in  the  raw  ore. 

Arsenic,  when  heated  in  air,  easily  oxidizes  into  white  arsenous 
oxide,  AS2O5,  and  is  easily  volatilized.     When  arsenical  ores  are  roasted, 

'  Trans.  A.  I.  M.  E.,  82-83. 


the  sulphur  and  arsenic  are  converted  into  arsenous  oxide  and  sulphur 
dioxide.  The  former  is  a  solid  at  ordinary  temperatures,  and  the  latter 
gaseous,  and  therefore  the  arsenous  oxide  is  deposited  as  a  sublimate  in 
the  cooler  portions  of  the  flues  and  dust  chambers,  through  which  the 
fumes  escape  from  the  furnace. 

In  roasting  cupriferous  pyritic  ores  there  will  usually  be  no  difficulty 
in  eliminating  from  75  to  80  per  cent,  of  the  arsenic. 

Roasting  Argentiferous  Cobalt -nickel  Arsenides  \ — The  ore  used  in  this 
investigation  was  chiefly  smalltite,  containing  689  oz.  silver  per  ton,  and 
56  per  cent,  arsenic.  The  object  of  the  investigation  was  to  ascertain 
(1)  the  temperature  at  which  the  arsenic  is  most  rapidly  expelled;  (2) 
the  thoroughness  with  which  it  is  expelled  by  prolonged  roasting  at  this 
temperature,  and  (3)  the  effect  of  adding  charcoal  near  the  end  or  at  the 
beginning  of  the  roast. 

It  was  found  that  15  per  cent,  of  arsenic  per  100  of  ore,  that  is,  27  per 
cent,  of  the  total  arsenic,  is  expelled  below  700°  C,  but  that  the  rest  of  the 
arsenic  is  not  expelled  until  the  temperature  reaches  about  840°  C.  when 
rapid  expulsion  sets  in.  By  rabbling  at  temperatures  above  840°  C,  the 
percentage  of  arsenic  can  be  further  reduced  by  about  34  per  cent.,  that 
is  down  to  17  per  cent,  in  the  ore,  from  the  original  56  per  cent;  in  this 
range  of  temperature  the  arsenic  is  removed  much  faster  than  at  lower 
temperatures.  Raising  the  temperature  quite  suddenly  to  800°  C.  does  no 
harm  as  the  ore  remains  porous.  The  addition  of  charcoal  either  at  the 
beginning  or  toward  the  end  of  the  roast  failed  to  increase  the  expulsion 
of  arsenic.  Finer  grinding  of  the  ore,  after  it  had  been  roasted  once, 
and  re-roasting  at  about  880°  C.  showed  no  further  expulsion  of  arsenic, 
due  to  fine  grinding. 

Antimony. — Antimony,  like  arsenic,  usually  occurs  as  the  sulphide. 
It  is  almost  universally  in  the  form  of  Stibnite  (SbjSa)  and  is  frequently 
associated  with  silver  and  gold,  and  quite  commonly  with  copper.  In  the 
small  quantities  in  which  it  usually  occurs  in  the  ores  of  these  metals,  it 
does  not  present  any  special  difficulties  in  roasting.  Care  must  be  used 
in  the  early  stages  of  the  operation.  If  an  attempt  were  made  to  remove 
the  antimony  by  rapid  oxidation,  there  would  be  danger  of  converting 
it  into  the  insoluble  antimonates  of  the  metals  in  the  ore.  This  would  be 
undesirable  for  some  of  the  chemical  processes,  while  for  others  it  might 
be  somewhat  serious.  In  the  early  stages  of  the  roasting  it  is  therefore 
necessary  to  employ  a  very  low  heat.  The  presence  of  steam,  largely 
supplied  by  the  burning  fuel  and  the  water  of  hydration  in  the  ore,  is 
found  to  be  useful  as  a  source  of  hydrogen,  which  removes  sulphur  as 
hydrogen  sulphide: 

SbjSg -I-3H2 -l-heat  =  3H2S -l-Sb. 
'  Bi-Monthhj  Bulletin  of  the  A.  I.  M.  E.,  .Jan.  1907. 


The  antimony  then  combines  with  oxygen  and  escapes  as  a  volatile 

When  the  temperature  of  the  roasted  ore  is  brought  to  about  350°  C. 
(662°  F.)  the  atmospheric  oxygen  converts  the  antimony  trisulphide  into 
antimonous  oxide  and  sulphur  dioxide.  Antimonic  acid  is  formed  in  the 
presence  of  the  oxides  of  the  other  metals,  and  combines  with  them  to 
form  antimonates.  Sulphates  of  antimony  are  not  formed  during  the 
roasting.  If  the  ore  contains  large  quantities  of  foreign  sulphides, 
which  on  being  roasted  would  form  sulphates,  antimonates  of  the  foreign 
metals  are  formed  instead  of  the  sulphates. 

Carbon,  such  as  coal  or  charcoal,  finely  ground  or  mixed  with  the  ore, 
has  been  used  to  break  up  the  antimonates  and  arsenates,  and  expel  the 
arsenic  and  antimony,  but  it  has  not  been  found  of  utility  enough  to  find 
a  permanent  place  in  practice. 

All  ores  of  copper,  gold,  and  silver,  containing  appreciable  quantities 
of  arsenic  and  antimony,  are  difficult,  if  not  impracticable,  to  treat  raw, 
by  any  of  the  solvent  processes.  By  roasting,  and  consequent  volatili- 
zation of  these  elements, .they  are  largely  eliminated  from  further  consid- 
eration, except,  perhaps  in  the  case  of  silver  ores  to  be  treated  by  sodium 
hyposulphite,  when  the  ores  may  contain -considerable  quantities  of  arsen- 
ates and  antimonates,  after  roasting. 

Antimonous  oxide  (SbjOg  or  Sb^OJ  resulting  from  the  roasting  of 
antimony  sulphide  (SbjSg)  is  insoluble  in  water,  but  is  soluble  in  hydro- 
chloric acid  and  in  alkalies. 

Bismuth. — Bismuth  is  one  of  the  most  injurious  alloys  of  copper.  It 
may  be  present  in  copper  ores  in  the  metallic  state,  or  in  sulphides, 
arsenides,  and  antimonides.  The  metal  and  sulphides  are  volatile  at  the 
roasting  temperature  but  much  less  readily  than  in  the  cases  of  arsenic 
and  antimony.  The  minerals  containing  bismuth  are  readily  oxidized 
to  fixed  compounds. 

In  the  incomplete  roasting  of  copper  ores,  arsenic,  antimony,  and 
bismuth  may  remain  in  the  roasted  product  in  the  same  combinations 
in  which  they  occurred  in  the  ores,  accompanying  the  fixed  compounds 
that  are  formed  during  the  roasting  operation. 

The  degree  of  elimination  of  these  impurities  in  roasting  varies 
necessarily  with  the  minerals  in  which  they  occur,  as  well  as  the  copper 
ore,  and  the  conditions  under  which  the  roasting  is  carried  on.  The 
following  analytical  data  by  Allan  Gibb'  shows  the  elimination  from  fairly 
typical  ores,  when  roasted  in  heaps  and  in  reverberatory  furnaces  for 
smelting,  which  does  not  represent  as  complete  a  roast  as  that  required 
for  the  wet  methods. 

'  Trans.  A.  I.  M.  E.,  Vol.  XXXIII. 







Roasted  ore 

Per  cent, 

Per  cent, 
Cu  =  100 

Per  cent, 
actual        1 


per  100 

of  Cu 

Per  cent, 
Cu  =  100 

Total  per- 
centage of 


.  1 



























14.68         ' 








0.045       ' 







0.015       1 

0.013       1 



No.  1  was  a  cupriferous  iron  pyrites  which  was  roasted  in  heaps  and 
subsequently  smelted  in  blast  furnaces. 

No.  2  was  a  dressed  ore  containing  the  copper  mostly  in  the  form  of 
copper  pyrites,  with  a  small  proportion  of  bornite  and  copper  glance. 
It  was  roasted  in  a  reverberatory  furnace. 

Nickel. — Nickel  is  quite  frequently  associated  with  copper  ores,  and 
when  it  so  occurs  in  paying  quantities  its  recovery  is  advisable. 

Nickel  usually  occurs  as  the 

Sulphide,  Millerite,  NiS, 

Arsenide,  Niccolite,  NiAs, 

Silicate,  Garnierite,  H  (NiMg)Si04,  H^O. 

When  the  sulphide  is  roasted,  the  nickel  is  oxidized  and  the  sulphur 
passes  off,  mostly  as  the  dioxide,  and  some  as  the  trioxide.  The  tri- 
oxide  produces  sulphuric  acid  and  forms  some  nickel  sulphate.  When 
the  sulphate  is  strongly  heated  the  nickel  is  converted  into  nickelous 
oxide  and  sulphur  trioxide  is  driven  off.  By  prolonged  roasting,  at  the 
proper  temperature,  nickelous  oxide,  NiO,  alone  may  be  obtained.  If 
imperfectly  roasted,  there  will  be  a  mixture  of  oxide,  ■  sulphate,  and 
unaltered  sulphide. 

If  a  mixture  of  nickel  and  iron  sulphides  is  carefully  roasted,  a  mix- 
ture of  nickelous  oxide  and  ferric  oxide  is  obtained.  As  sulphate  of 
nickel  is  a  very  stable  compound,  the  roasting  may  be  so  conducted, 
that  the  greater  part  of  the  nickel  is  obtained  as  sulphate,  while  the  iron 
will  be  in  the  condition  of  ferric  oxide. 


By  roasting  nickel  and  copper  sulphides  in  the  same  way,  it  is  possible 
to  get  nickelous  oxide  and  cupric  oxide,  or  a  mixture  of  oxides  and  sul- 
phates. As  nickel  sulphate  is  stable  at  a  higher  temperature  than 
copper  sulphate,  the  nickel  may  be  roasted  to  the  sulphate  and  the  copper 
to  the  oxide. 

If  nickel,  iron,  and  copper  sulphides  are  all  roasted  together,  the 
nickel  may  be  in  the  condition  of  sulphate  and  the  other  two  metals  as 
ferric  and  cupric  oxides. 

If  nickel  arsenide  is  roasted,  the  arsenic  forms  arsenous  oxide,  and 
the  nickel  sesquioxide.  Part  of  the  arsenous  oxide  escapes  unaltered, 
part  is  further  oxidized  to  arsenic  oxide,  and  this  combines  with  the 
nickelous  oxide  to  form  an  arsenate.  Nickel  arsenate  is  not  decomposed 
when  heated  alone,  so  the  result  is  basic  nickel  arsenate. 

Copper  ores  containing  nickel  usually  contain  also  magnetic  and  iron 
pyrites,  and  often  arsenic  and  antimony  compounds  as  well  as  silicates, 
quartz,  and  earthy  matter.  In  the  roasting,  the  arsenic  and  antimony  are 
mostly  driven  off,  the  sulphur  partly  escapes  as  dioxide,  and  is  partly 
converted  into  trioxide  by  contact  with  the  red-hot  masses  of  ore  and 
furnace  walls.  Iron,  copper,  and  nickel  oxides  combine  with  this  trioxide 
to  form  sulphates.  As  the  roasting  proceeds,  and  the  temperature  is 
raised,  the  sulphates  are  again  decomposed  into  oxides  and  sulphur 
trioxide,  or  sulphur  dioxide  and  oxygen.  Iron  sulphate  is  first  decom- 
posed, next  the  copper,  and  lastly  the  nickel  compound.  If  the  roasting 
were  continued  at  the  proper  temperature,  the  product  would  be  a 
mixture  of  ferric  oxide,  cupric  and  cuprous  oxides,  and  nickelous  oxide. 

Nickel  sulphate  is  readily  and  abundantly  soluble  in  water.  The 
oxide  is  soluble  in  mineral  acids,  especially  dilute  hydrochloric  acid, 
when  warmed.  The  chloride  is  soluble  in  water,  but  not  as  readily  as 
the  sulphate.  The  mineral  garnierite  is  soluble  in  sulphuric  and  hydro- 
chloric acids,  but  with  some  difficulty. 

Calcium  (Lime). — The  compounds  of  calcium,  on  account  of  their 
prevalence  and  positive  action  on  almost  all  of  the  chemical  solvents  used 
in  the  hydrometallurgical  processes,  are  among  the  most  important  to  be 
considered.  In  the  alkali  process,  calcium  compounds  are  not  particu- 
larly harmful,  and  frequently,  as  in  the  case  where  calcium  is  combined  with 
oxygen  to  form  lime,  it  imparts  a  desired  alkalinity  before  applying  the 
solvent.  In  the  acid  processes,  like  chlorination  or  the  treatment  of  copper 
ores  with  dilute  acids,  the  amount  of  calcium  in  the  ore  and  the  way  it  is 
combined  will  usually  be  the  most  important  factor  in  determining  the 
applicability  of  the  process  and,  to  a  large  extent,  indicate  its  success  or 

There  would  be  no  difficulty  in  treating  most  of  the  copper,  gold  and 
silver  ores  successfully,  by  the  acid  processes,  if  it  were  not  for  a  few 
interfering  elements,  and  of  all  the  interfering  elements,  the  presence  of 


lime  in  large  quantities  presents  the  most  common  and  the  most  difficult 
problem.  Fortunately,  most  of  the  ores  of  copper  and  the  precious 
metals  do  not  contain  enough  lime  to  seriously  interfere  with  the  treat- 
ment. The  vast  majority  of  all  metalliferous  deposits  have  quartz  as 
the  matrix,  and  usually  the  lime  is  not  found  in  quantities  sufficient  to 
make  an  acid  treatment  prohibitive,  if  the  ore  is  otherwise  suited  to  the 
process,  particularly  if  the  ore  is  amenable  to  preliminary  concentration. 
Calcium  usually  occurs  associated  with  copper,  gold  and  silver  ores 
in  the  form  of 

The  Carbonate  (CaCOg),  Calcite  (Limestone). 
The  Fluoride  (CaF^),  Fluorite. 
The  Sulphate  (CaSOJ,  Gypsum. 

The  carbonate  is  not  readily  attacked  by  chlorine,  but  is  immediately 
decomposed  by  acids.  Roasting  converts  the  carbonate  into  the  oxide 
and  carbon  dioxide: 

CaCOj  +  Heat  =  CaO  +  CO^. 

When  cold,  the  oxide  (lime)  does  not  absorb  chlorine,  but  at  a  red  heat, 
in  the  presence  of  chlorine,  it  forms  calcium  chloride  with  the  evolution 
of  oxygen: 

CaO  +  2Cl  =  CaCl2  +  0. 

If  the  ore  contains  considerable  sulphur,  the  sulphur  trioxide  released 
during  the  roasting  combines  to  a  greater  or  less  extent  with  the  lime  to 
form  sulphate: 

CaO  +  S03  =  CaSO, 

which  is  practically  unaffected  by  all  the  acids,  only  very  slightly  acted 
upon  by  chlorine,  and  remains  neutral  to  cyanide  or  sodium  hyposulphite. 
It  is  almost  insoluble  in  water;  one  part  of  calcium  sulphate  requires  432 
parts  of  water  for  its  solution.  Its  solubility  is  increased  by  the  presence 
of  alkaline  chlorides  and  free  hydrochloric  acid. 

It  is  desirable,  therefore,  that  ores  containing  considerable  lime  should 
be  mixed  with  ores  containing  considerable  sulphur  before  roasting. 
Some  of  the  calcium,  however,  will  unavoidably  remain  as  oxide  after 
roasting,  which,  when  coming  in  contact  with  water  in  the  subsequent 
chemical  treatment,  is  converted  into  the  hydroxide  (slacked  lime)  in 
which  form  it  is  desirable  in  the  alkali  processes,  but  is  readily  attacked 
by  chlorine  and  the  acids. 

The  lime,  when  coming  in  contact  with  sulphuric  acid,  as  in  the  sul- 
phuric acid  copper  processes,  and  the  barrel  chlorination  process  where 
chlorine  is  generated  from  bleach  and  acid,  is  converted  into  calcium 

Ca(OH)2  +  H2S04  =  CaSO,  +  2H20, 


which  accounts  for  much  of  the  excess  of  acid  sometimes  used  in  treating 
ores  containing  considerable  lime. 

.  In  order  to  economize  acid,  it  is  desirable  to  convert  as  much  as 
possible  of  the  lime  into  the  sulphate,  by  judicious  roasting,  if  the  ore  is 
a  sulphide. 

If  the  chlorinating  is  done  by  the  "Plattner"  or  by  the  "Percolation" 
processes,  in  which  acids  are  not  ordinarily  used,  then  the  lime,  instead 
of  combining  with  the  excess  of  acid,  will  combine  with  the  chlorine: 

2Ca(OH)2  +  4Cl  =  CaCl2  +  Ca(C10)2  +  2H20, 

forming  the  chloride  and  hypochlorite,  as  in  the  manufacture  of  bleaching 
powder.  Since  chlorine  acts  more  readily  on  lime  than  on  copper,  gold, 
and  silver,  in  the  ore,  sufficient  chlorine  must  be  provided  to  chlorinate  ' 
the  lime  and  have  an  excess  after  all  other  base  elements  have  been 
satisfied.  To  avoid  the  large  consumption  of  chlorine  when  it  is  applied 
directly  as  gas  in  ores  containing  much  lime,  the  ore  is  frequently  roasted 
with  salt;  in  this  way  the  lime  is  converted  into  chloride  in  the  furnace  and 
is  no  longer  harmful. 

Calcium,  in  the  form  of  fluorite,  is  a  common  associate  of  copper,  gold 
and  silver  ores.  It  occurs  abundantly  in  Cripple  Creek,  intimately  associ- 
ated with  calcite.  Fluorite  is  peculiarly  a  constituent  of  metalliferous 
veins.     In  minute  quantities  it  is  widely  diffused. 

Fluorite  is  unaffected  by  chlorine,  cyanide,  or  dilute  acids.  Hot 
concentrated  sulphuric  acid  decomposes  it.  By  roasting,  the  fluorite  is 
converted  into  the  oxide,  as  in  the  case  of  carbonate: 

CaF2  +  H20  =  CaO+2HF. 

"  The  fluorine  probably  combines  with  the  moisture  of  the  air,  and 
water  combined  in  the  ore  as  hydrate,  to  form  hydrofluoric  acid. 

Mixed  with  silica  and  sulphur,  as  the  fluorite  usually  is  in  metallifer- 
ous ores,  the  sulphuric  acid  formed  in  the  roasting  converts  some  of 
the  calcium  into  the  sulphate: 

CaF2-FH2SO,  =  CaSO,  +  2HF. 
Car2-t-2H2SO,-fSi02  =  CaS04-h2H20-FSiF2. 

The  principal  point  of  interest,  so  far  as  roasting  for  the  subsequent 
chemical  treatment  is  concerned,  is,  like  the  carbonate,  the  fluoride  is 
converted  into  lime,  and  that  in  the  presence  of  sulphur  it  is  converted 
into  the  sulphate. 

A  specimen  sample  of  Cripple  Creek  ore,  composed  largely  of  fluorite, 
after  roasting  had  a  white  appearance,  and  analysis  showed  39.25  per 
cent.  CaO. 

Calcium  sulphate  is  largely  associated  with  copper,  gold,  and  silver 
ores,  in  the  form  of  gypsum  or  anhydrite.     It  also  occurs  largely  as  the 


result  of  the  decomposition  of  pyritic  ores  acting  on  the  calcium  carbonate. 
Pyritic  ore  is  oxidized  by  the  action  of  water  and  air,  forming  ferrous 
sulphate  and  sulphuric  acid. 

FeS2  +  H20  +  0,=FeSO,  +  H2SO,. 

The  sulphuric  acid  then  acts  on  the  carbonate,  forming  calcium  sulphate 
and  water: 

CaC03  +  H2SO,  =  CaSO,  +  H20. 

Calcium,  therefore,  in  oxidized  ores  is  largely  in  the  form  of  sulphate, 
and  is  not  particularly  injurious  in  any  of  the  chemical  processes,  or  in 
the  roasting  operation.  The  sulphate,  once  formed,  can  only  be  converted 
into  the  oxide  by  the  most  intense  heat  procurable.  Such  a  heat  is  never 
realized  in  a  roasting  furnace. 

Gypsum  gives  off  its  water  of  hydration  at  200  to  250°  C.  (392  to 
482°  F.) .  The  dehydrated  gypsum  melts  at  a  red  heat  mthout  decomposi- 
tion. On  coming  in  contact  with  water,  the  dehydrated  calcium  sulphate 
again  takes  up  its  water  of  hydration,  just  as  in  the  case  of  ordinary 
plaster  of  Paris.  In  doing  this,  if  the  ore  contains  considerable  sulphate, 
it  sometimes  happens  that  the  ore  during  leaching  sets  so  hard  that  picks 
have  to  be  used     to  remove  it  from  the  vats. 

An  analysis  made  on  unoxidized  ore  from  Cripple  Creek,  showed: 

Calcium  Sulphate  (gypsum,  CaS04-|-2H20),  0.83  per  cent. 
Calcium  Fluoride  (fluorite,  CaFj),  0.78  per  cent. 

The  amount  of  lime  in  an  ore  which  may  be  fatal  to  chlorination  or 
to  an  acid  treatment  depends  largely  on  other  conditions.  Ordinarily 
from  5  to  6  per  cent,  is  the  limit.  In  Cripple  Creek  ores  the  lime  varies 
from  1.5  to  2.5  per  cent.,  although  in  some  mines  it  is  much  higher.  The 
Potsdam  ores  of  the  Black  Hills,  which  have  been  successfully  chlorin- 
ated, contain  as  much  as  8  per  cent.  CaO. 

That  only  a  small  portion  of  the  lime  in  roasted  ore  combines  with 
chlorine  or  the  acids  is  evident  from  the  treatment  of  800  to  1000  tons 
daily  of  Cripple  Creek  ores  by  the  barrel  chlorination  process,  where  it 
may  be  assumed  that  the  ore  averages  2per  cent,  lime,  or  40  lb.  per  ton. 
The  average  chemical  charge  may  be  assumed  to  be  15  lb.  of  bleach  and 
30  lb.  of  sulphuric  .acid.  Theoretically,  it  takes  6  parts  of  acid  to  com- 
bine with  7  parts  of  bleach,  but  in  practice,  owing  to  the  impurities  of 
the  bleach  and  acid,  equal  parts  of  each  are  required.  Of  the  30  lb. of  acid, 
therefore,  per  ton  of  ore  charged  into  the  barrel,  15  lb.  are  consumed  in 
reacting  with  the  bleach  to  generate  chlorine.  The  solutions  issuing  from 
the  barrels  after  treatment  are  always  strongly  acid,  so  that  much 
acid  remains  unconsumed,  and  some  is  also  consumed  in  reacting  with 
other  base  elements.  It  is  safe  to  say,  therefore,  that  only  from  5  to  10  lb. 


of  the  acid  actually  combines  with  the  calcium  or  lime  in  the  ore.  But 
if  the  calcium  in  the  ore  were  all  present  as  lime,  that  is  40  lb.  CaO,  it 
would  take  at  least  70  lb.  of  acid  to  neutralize  this  lime,  instead  of  only 
5  or  10  lb.  actually  required  in  practice.  Some  of  the  Cripple  Creek  ores 
are  chlorinated  with  only  10  lb.  of  bleach  and  15  lb.  of  acid,  which  makes 
the  acid  consumed  considerably  less.  From  this  it  will  be  seen  that  the 
injurious  effect  of  the  lime  in  ore  in  an  acid  process  depends  largely  on  its 
chemical  combination,  and  that  much  of  the  lime  in  sulphide  ores  may 
be  converted  into  a  comparatively  harmless  condition  by  roasting. 

Magnesium. — Copper,  gold  and  silver  ores  frequently  contain  small 
quantities  of  magnesium,  but  usually  not  in  sufficient  quantity  to  seri- 
ously interfere  with  any  operation  in  the  hydrometallurgical  processes. 
For  all  practical  purposes  of  hydrometallurgy  it  may  be  considered  as 
equivalent  to  its  analogous  element,  calcium.  Magnesium  usually  occurs 
combined  with  calcium  as: 

The  Carbonate,  (CaMg)C03,  Dolomite, 
The  Sulphate,  MgSO^.H^O,  Kieserite, 
The  Silicate,  H^MgjSiOg,  Serpentine, 
The  Silicate,  H2Mg3(Si03)4,  Talc. 

In  oxidized  ores  the  magnesium  is  largely  in  the  form  of  carbonate  and 
silicate.  It  may  also  be  present  as  sulphate,  formed  by  the  decomposi- 
tion of  pyrites,  as  the  corresponding  calcium  sulphate.  The  magnesium 
sulphate,  kieserite,  is  very  slowly  soluble  in  water — about  like  gypsum. 
The  hydrous  sulphate  epsomite  (MgSO^jXHjO)  is  readily  soluble.  In 
roasting,  this  water  of  hydration  is  driven  off.  Much  of  the  magnesium 
sulphate  formed  in  the  oxidation  of  pyrites  in  mineralized  veins  is  carried 
away  in  solution. 

In  roasting  sulphide  ores,  the  magnesium  carbonate  is  partly  con- 
verted into  the  oxide  and  partly  into  the  sulphate.  The  oxide,  like  the 
corresponding  calcium  oxide,  is  practically  insoluble  in  water.  It 
reacts  readily  with  chlorine,  bromine,  hydrochloric  and  sulphuric  acids,  to 
form  the  chloride,  bromide,  and  sulphate.  Magnesium  chloride  is  very 
soluble  in  water — 100  parts  of  water  will  dissolve  about  52  parts  of 
magnesium  chloride  at  ordinary  temperatures. 

Magnesium  sulphate  is  practically  unaffected  by  any  of  the  chem- 
ical solvents.  All  the  harmful  hydrous  sulphates  may  be  converted 
into  the  harmless  anhydrous  Sulphate  by  roasting.  As  in  the  case  of 
calcium,  therefore,  ore  containing  magnesium  should  be  roasted  with  a 
view  of  converting  as  much  of  it  as  possible  into  the  form  of  sulphate. 

Talc  is  insoluble  in  acids  both  before  and  after  ignition.  Roasting 
greatly  improves  the  talc  for  subsequent  treatment  by  the  wet  methods, 
especially  in  the  leaching  or  filtering  qualities  of  the  ores  containing  it. 


Manganese. — Manganese  is  one  of  the  most  deleterious  substances  in 
the  extraction  of  metals  by  wet  methods.  It  affects  injuriously  the  acids, 
the  halogens,  and  cyanide.  Fortunately  it  does  not  frequently  occur 
in  ores  of  copper  and  the  precious  metals  in  quantities  so  great  as  to  be 

Manganese  almost  universally  occurs  as  the  oxide;  sometimes  as  the 
sulphide  and  silicate.  After  roasting  it  is  always  in  the  form  of  oxide, 
and  roasting  does  not  materially  lessen  its  injurious  effects  on  the  solvent. 
Manganese  is  readily  soluble  in  acids  and  difficult  to  eliminate  from  the 
solvent.     Its  principal  injurious  effect  is  in  the  consumption  of  chemicals. 

Aluminum. — Aluminum,  as  it  occurs  in  copper,  gold,  and  silver  ores, 
affects  the  chemical  processes  somewhat  injuriously.  Its  mineralogical 
combinations  are  numerous  and  varied.  It  may  occur  as  the  oxide, 
hydroxide,  sulphate,  or  silicate.  It  usually  occurs  as  the  silicate,  more 
or  less  intimately  associated  with  calcium,  magnesium,  iron,  and  the 
alkali  metals. 

Roasting  converts  some  of  the  aluminum  compounds  into  aluminum 
oxide  (AI2O3),  which  is  infusible  at  all  temperatures  ever  attained  in  a 
roasting  furnace.  It  is  not  decomposed  by  heat  alone.  It  is  not  decom- 
posed by  chlorine  at  any  temperature.  Anhydrous  aluminum  oxide  is 
perfectly  insoluble  in  water.  After  strong  ignition,  it  is  likewise  insoluble 
in  most  acids.  The  lower  the  temperature  at  which  aluminum  oxide  is 
heated,  the  more  soluble  it  is  in  the  acids  and  alkalis. 

All  the  silicates  of  aluminum  are  insoluble  in  water,  with  the  exception 
of  the  alkali  salts,  and  these  are  soluble  only  when  the  ratio  of  the  base  to 
the  acid  is  above  a  certain  limit.  Many  of  the  silicates  are  decomposed 
by  dilute  sulphuric  and  hydrochloric  acids.  Chlorine,  bromine,  and 
potassium  cyanide  react  very  slowly. 

Aluminum  sulphate,  AljCSOJg,  when  heated  to  redness,  is  converted 
into  the  oxide.  The  sulphate  is  very  soluble  in  water.  Chlorine  reacts 
very  slowly  with  it.  The  basic  sulphate,  AljOg,  SO3,  lOHjO,  is  insoluble 
in  water,  but  soluble  in  sulphuric  and  hydrochloric  acids. 

When  ores  containing  considerable  aluminum  are  properly  roasted, 
and  a  sample  filtered  with  water,  it  will  be  found  on  testing  that  there  is 
no  soluble  aluminum  in  the  ore.  If  the  sample  is  then  filtered  with  di- 
lute sulphuric  or  hydrochloric  acid,  some  aluminum  will  be  dissolved. 
If  the  sample  is  treated  with  chlorine,  bromine,  or  potassium  cyanide, 
only  traces  will  be  found  in  the  solution. 

The  compounds  of  aluminum  are  so  numerous,  varied  and  compli- 
cated that  it  is  difficult,  if  not  impossible,  to  determine  their  exact 
composition  either  in  the  raw  or  roasted  ore.  The  only  alternative  seems 
to  be  to  resort  to  direct  tests  with  the  chemical  solvents.  If  acids  are 
used  in  the  chemical  treatment  of  the  ore,  some  of  the  consumption  of 
the  acid  is  due  to  combining  with  aluminum.     Beyond  the  slightly 


increased  cost  of  treatment,  due  largely  to  increased  consumption  of  acid, 
no  great  inj  ury  to  its  presence  in  the  ore  is  apparent.  Its  presence,  even 
in  large  quantities,  is  not  fatal,  or  even  serious,  to  any  chemical  process. 
Cripple  Creek  ores,  which  are  very  successfully  treated  after  roasting,  by 
cyanidation,  and  by  chlorination  with  or  without  the  us&  of  acid,  fre- 
quently contain  as  high  as  20  per  cent,  alumina  (AI2O3),  and  the  average 
is  about  18  per  cent.  Copper  ores  at  Clefton,  Arizona,  containing  16  per 
cent,  alumina  have  been  successfully  leached  for  many  years  with  sul- 
phuric acid. 

Usually  copper,  gold,  and  silver  ores  do  not  contain  more  than  several 
per  cent,  alumina;  frequently  it  is  less  than  1  per  cent.  Whatever  the 
condition  of  the  aluminum  in  the  raw  ore,  where  it  may  be  injurious,  the 
tendency  in  roasting  is  to  convert  it  into  the  harmless  aluminum  oxide. 
The  higher  the  temperature  at  which  the  ore  is  roasted,  the  less  difficulty 
will  result  due  to  the  presence  of  aluminum,  but  the  ultimate  temperature 
of  roasting  ores  containing  much  aluminum  will  depend  on  the  other, 
more  or  less  fusible,  constituents. 

The  hydrate  of  aluminum  occurs  mineralogically  as  Gibbsite;  it  is 
easily  dissolved  by  acids.  The  monohydrate  occurs  native  as  diaspora; 
it  gives  up  its  water  of  hydration  at  360°  C.  (680°  F.). 

Clay. — This  is  the  term  applied  to  hydrous  silicates  of  aluminum, 
produced  for  the  most  part  by  the  decomposition  of  feldspar  rocks,  and 
generally  mixed  with  other  substances,  chiefly  lime,  magnesia,  and  oxide 
of  iron.  Clay  is  frequently  a  constituent  of  ores,  usually  occurring  as 
"Gouge"  matter  in  the  vein. 

As  a  rule  clays  contain  from  45  to  60  per  cent,  silica;  from  20  to  30 
per  cent,  alumina;  from  0.5  to  3  per  cent,  lime;  from  0.5  to  3  per  cent, 
magnesia,  small  quantities  of  iron,  and  about  19  per  cent,  water.  Clays 
always  contain  a  hydrous  compound  of  alumina  and  silica,  which  is  able 
to  give  up  the  alumina  contained  by  it  as  a  base  to  sulphuric  acid. 

Clays  are  very  much  improved  by  roasting,  both  as  to  filtration  and 
chemical  consumption. 

Barium  frequently  occurs  associated  with  copper,  gold,  and  silver 
ores  in  small  quantities.  It  is  usually  in  the  form  of  sulphate,  Barite 
(heavy  spar,  BaSOJ.  Sometimes  it  occurs  as  the  carbonate,  Witherite 

If  the  carbonate  is  heated  in  an  atmosphere  free  from  sulphur,  the 
barium  oxide,  BaO,  will  be  produced,  which  reacts  with  the  halogens  and 
the  acids.  The  temperature  required  for  the  decomposition  of  the 
carbonate  by  heat  alone  is  very  high.  In  the  presence  of  sulphur,  the 
carbonate  is  converted  into  the  sulphate. 

Barium  sulphate  is  practically  unaffected  by  any  operation  of  the 
chemical  processes.  Any  heat  obtainable  in  a  roasting  furnace  does  not 
decompose  it.     It  is  insoluble  in  water  and  in  acids. 


Alkali  Metals. — The  alkali  metals,  sodium,  and  potassium,  are  fre- 
quently found  in  considerable  quantities  associated  with  ores.  They 
usually  occur  as  the  feldspars  or  hornblende,  and  as  such  are  unaffected 
by  roasting  or  any  of  the  chemicals  used  in  the  solvent  processes. 

Chlorine,  Bromine. — Chlorine  and  bromine  are  sometimes  found  in 
copper,  gold,  and  silver  ores,  and  when  they  do  so  occur  are  of  considerable 
metallurgical  importance  in  roasting.  The  compound  which  is  most 
common  is  the  silver  chloride,  cerargyrite  (AgCl) .  The  minerals  embo- 
lite,  Ag(ClBr),  and  bromyrite,  AgBr,  occur  occasionally,  and  in  roasting 
may  be  considered  the  same  as  cyrargyrite.  Chlorine  also  occurs  quite 
frequently  in  combination  with  lead. 

The  surface  ores  of  Tonapah,  Nevada,  show  much  of  the  silver 
combined  with  chlorine — frequently  as  much  as  20  per  cent.  As  depth  is 
attained,  the  silver  chloride  gradually  merges  into  the  sulphide,  although 
the  chlorides  appear  never  to  be  entirely  absent. 

The  principal  point  of  importance  in  connection  with  the  roasting  of 
silver  chloride,  is  the  danger  of  volatilization,  even  with  an  oxidizing 
roast.  In  making  exhaustive  tests  in  Denver,  on  a  working  scale,  on 
some  of  the  Tonapah  ores,  it  was  found  that  the  volatilization,  with  an 
oxidizing  roast,  was  about  the  same  as  with  a  chloridizing  roast,  but  in  no 
case  was  the  volatilization  serious.  If  volatilization  is  known  to  take 
place  in  an  oxidizing  roast,  chlorides  in  the  ore  may  be  suspected. 

To  ascertain  the  amount  of  silver  chloride  in  the  ore,  leach,  or  treat 
a  sample  with  sodium  hyposulphite  (sodium  thiosulphate)  and  compare 
the  hypo  tails  with  that  of  the  original  ore.  Also  test  for  chlorine  with 
silver  nitrate. 

Loss  of  Weight  in  Roasting. — There  is  always  some  loss  of  weight  in 
ore  due  to  roasting.  The  loss  is  usually  largest  in  pyritic  ores  and  in 
pyritic  concentrates,  but  it  may  also  be  considerable  in  ores  which  are 
oxidized  and  highly  silicious.  In  sulphide  ore  the  loss  is  due  mostly  to 
the  expulsion  of  the  sulphur;  in  oxidized  ores  it  is  mostly  due  to  driving 
off  the  water  of  hydration.  The  water  so  combined,  in  many  ores  may 
be  quite  large.  Iron  in  oxidized  unroasted  ores  is  almost  always  in  the 
form  of  ferric  hydrate  (limonite),  2re03,  SHsO,  which  contains  14.4  per 
cent,  water,  all  of  which  is  driven  off  in  roasting.  Similarly  othei'  sub- 
stances give  up  their  water  of  hydration,  and  some  of  the  elements  are 
eliminated  by  volatilization. 

The  loss  of  weight  in  sulphide  ores  is  represented  by  the  substitution 
of  oxygen  for  sulphur.     From  the  equation 

4FeS2  +  1102  =  2Fe203+8S02 

the  loss  of  weight  of  pyrites  can  readily  be  calculated  that  3  parts  of 
FeS2  =  2  parts  Fefis,  but  the  matter  is  usually  not  so  simple  as  this,  owing 
to  other  constituents  in  the  ore  and  the  manner  in  which  the  remaining 


sulphur  is  combined.  If,  for  example,  there  is  galena  (PbS)  in  the  ore 
and  is  oxidized  to  sulphate  (PbSO^),  there  has  been  an  actual  gain  of 
weight  of  4  atoms  of  oxygen,  or  27  per  cent. 

The  loss  of  weight  can  readily  be  ascertained  from  the  difference  in 
weight  between  the  raw  and  roasted,  as  it  is  charged  and  withdrawn 
from  the  furnace.  This  method  is  expensive  and  not  quite  accurate 
because  the  dust  loss  cannot  usually  be  taken  into  consideration.  The 
loss  of  weight  is  best  and  most  conveniently  obtained  by  direct  experi- 
ment. This  is  done  by  weighing  a  small  average  sample  of  the  ore,  then 
thoroughly  drying  it;  weighing  it  again,  and  then  roasting  it  in  a  roasting 
dish,  in  a  muffle,  to  the  same  extent  as  the  ore  is  roasted  in  the  mUl.  A 
sulphur  determination  will  show  this.  From  the  differences  in  weight 
between  the  raw  ore,  the  dried  ore,  and  the  roasted  ore,  the  loss  due  to 
moisture  and  the  loss  due  to  roasting  can  easUy  and  accurately  be 

A  ton  of  roasted  pyritic  concentrates  will  occupy  about  24  1/2  cu.  ft. 
This  is  derived  from  2800  lb.  of  raw  ore,  which  will  occupy  about  23  2/3 
cu.  ft.  per  ton.  A  ton  of  the  concentrates  after  roasting  will  weigh  from 
1450  to  1700  lb.,  and  will  occupy  about  17  1/2  cu.  ft.  The  loss  of  weight 
in  Cripple  Creek  ores,  due  to  roasting,  is  usually  from  5  to  7  per  cent., 
based  on  the  control  samples.  Of  this  loss,  about  2  per  cent,  is  for 
moisture,  and  from  3  to  4  per  cent,  dust  and  volatilization  loss.  Of  the 
volatilization  loss  about  1  per  cent,  is  accounted  for  by  the  elimination 
of  the  greater  portion  of  the  sulphur.  The  accountable  dust  loss  is 
about  2  per  cent.,  and  the  unaccountable  loss  amounts  to  about  1  per  cent. 
Much  of  this  unaccountable  loss  is  due  to  unsettled  dust  going  out  of  the 
furnace  stacks,  and  some  also  due  to  unrecovered  dust  in  crushing  and 
roasting  other  than  flue  dust. 

At  Butte,  in  roasting  copper  concentrates,  containing  35  per  cent, 
sulphur  down  to  7  per  cent,  sulphur,  the  loss  of  weight,  including  flue  dust, 
is  about  20  per  cent. 

In  roasting  Black  Hills  ore,  containing  11  per  cent,  sulphur,  down  to 
0.08  per  cent.,  there  was  a  loss  in  weight  of  21  per  cent.,  even  though  the 
ore  was  apparently  thoroughly  dry.  This  ore  was  very  talcy,  and  the 
great  loss  was  evidently  due  principaUy  to  the  water  of  hydration. 


Object  of  Chloridizing  Roasting.— Most  of  the  chlorides,  at  elevated 
temperatures  and  in  the  presence  of  sulphides  or  sulphates,  have  the 
power  of  converting  copper  and  silver  into  their  respective  chlorides,  and, 
to  some  extent,  the  gold  also.  To  roast  in  the  presence  of  chlorides, 
usually  sodium  chloride  (common  salt),  is  known  as  "Chloridizing 
Roasting."  The  term  "chloridizing"  is  limited  to  the  production  of 
chlorides  by  the  interchanging  of  chlorine  from  its  chloride  combinations, 
usually  at  elevated  temperatures;  while  the  term  "chlorinating"  is 
limited  to  the  production  of  chlorides,  usually  in  the  wet  way,  by  the 
application  of  free  chlorine. 

The  objects  of  chloridizing  roasting  are: 

1.  In  copper  ores,  or  in  gold  and  silver  ores  containing  copper,  to 
convert  the  copper  into  chlorides,  which  will  not  react  with  chlorine  or 
the  acids,  but  which  are  directly  soluble  in  water  or  in  chloride 

2.  In  silver  ores,  or  in  gold  and  copper  ores  containing  silver,  to  con- 
vert the  insoluble  metallic  silver  or  its  insoluble  compounds,  into  the  more 
soluble  silver  chloride. 

3.  In  any  ore,  to  convert  the  harmful  elements  into  less  harmful 

4.  To  assist  in  a  more  efficient  oxidizing  action  than  is  possible  under 
the  same  conditions,  in  ordinary  oxidizing  roasting. 

Metallic  silver  is  not  readily  soluble  in  any  of  the  commercial  chemical 
solvents.  The  silver  chloride  is  readily  soluble,  either  in  chloride  solu- 
tions, sodium  or  calcium  hyposulphite,  or  in  potassium  or  sodium 
cyanide.  If,  therefore,  a  high  percentage  of  the  silver  can  be  converted 
into  the  chloride,  a  quick  and  correspondingly  high  percentage  of  the 
silver  can  be  extracted. 

If  the  ore  contains  copper,  or  if  a  copper  ore  is  treated  by  a  chloride 
process,  it  is  frequently  desirable  to  convert  the  copper  in  the  ore  into 
the  soluble  cupric  chloride,  so  as  to  save  acid,  if  an  acid  process  is  used. 
It  may  be  cheaper  to  convert  the  copper  into  chlorides  at  the  expense  of 
a  cheap  material,  such  as  salt,  than  to  let  the  oxides  react  with  the  more 
expensive  acids.  Chloridizing  roasting  is  largely  used  in  the  extraction 
of  copper  from  its  ores. 



Most  of  the  chlorides  are  soluble  in  water;  if  desired,  many  of  the 
objectionable  elements  in  the  ore  may  be  removed  by  a  preliminary 
washing,    after   roasting,    and   before   applying  the   chemical   solvent. 

Of  the  metallic  sulphides  usually  associated  with  copper,  gold  and 
silver  ores,  those  of  iron,  copper,  lead,  and  zinc  are  the  most  common. 
Of  these,  only  the  iron  and  copper  sulphides,  are  available  to  react  with 
the  salt;  while  those  of  lead  and  zinc  remain  quite  indifferent. 

Adaptability  of  the  Various  Ores  to  Chloridizing  Roasting. — Ottokar 
Hofmann'  aptly  classifies  the  adaptability  of  the  various  ores  to  chloridiz- 
ing roasting,  as  follows: 

1.  Those  like  iron  and  copper  pyrites,  gray  copper  ore,  and  silver 
copper  glance,  which  in  roasting  form  sulphates,  and  decompose  salt, 
liberating  chlorine. 

2.  Those  like  galena  and  zinc  blende,  which  form  sulphates  remaining 
indifferent  to  salt. 

3.  Antimonial  and  arsenical  silver  minerals  which  form  antimonates 
and  arsenates  of  silver. 

The  gangue  remains  indifferent,  like  quartz  or  porphyry,  or  it  takes 
an  active  part,  like  limestone,  talc,  spar,  manganese,  and  minerals  con- 
taining magnesia. 

If  ore  consists  of  minerals  of  the  first  group  together  with  an  indifferent 
gangue,  chloridizing  roasting  offers  no  difficulty  and  a  high  chloridization 
can  be  obtained  without  much  loss  of  silver  by  volatilization  and  no 
special  skill  is  required  in  the  roasting;  neither  does  it  matter  if  the  salt 
is  added  to  the  charge  before  entering  the  furnace  or  after  it  has  been 
subjected  to  partial  oxidizing  roasting. 

The  process  of  chloridizing  roasting  becomes  more  difficult  if  one  or 
both  of  the  minerals  of  the  second  class  are  present  in  large  quantities, 
even  if  associated  with  an  indifferent  gangue.  With  such  ores  the  time 
of  adding  the  salt  becomes  very  important.  If  added  before  the  charge 
enters  the  furnace  a  very  inferior  chloridization  is  obtained,  as  is  also  the 
case  if  the  salt  is  added  before  the  oxidizing  period  has  sufficiently 
advanced.  'Moreover,  the  temperature  and  air  supply  require  much 

The  roasting  is  still  more  difficult  if  all  the  classes  of  ore  are  repre- 
sented in  connection  with  a  gangue  like  limestone  which  takes  an  active 
and  injurious  part  in  the  operation. 

Chemistry  of  Chloridizing  Roasting. — The  sulphides  in  the  ore, 
mostly  relied  upon  for  chloridization,  are  those  of  iron  and  copper.  The 
sulphates  of  these  metals,  formed  during  the  roasting,  react  with  the 
salt  to  form  sodium  sulphate  and  the  chlorides  of  the  metals.  Some 
hydrochloric  acid  and  chlorine  are  formed  at  the  same  time,  largely  due 

'  Mineral  Industry,  1896. 


to  the  action  of  the  sulphur  trioxide  and  sulphuric  acid.     The  following 
reactions,  represent  in  a  general  way  the  chloridizing  action: 

2NaCl+  FeSO,  =  Na2S04+FeCl2. 
2NaCl+  CuS0,  =  Na,S0,  +  CuCl2. 
2NaCl+     2S03  =  Na3SOi  +  2Cl  +  S02. 
2NaCl+  H2SO,  =  Na2SO,  +  2HCl. 

The  chlorine  and  chlorides  thus  formed  react  with  the  silver  and  silver 
sulphate  to  form  the  silver  chloride: 

FeCl2+  AgS0,  =  2AgCl+FeS0,. 

CuCl,+  AgS0,  =  2AgCl  +  CuS0,. 
2NaCl  +  AgSO,  =  2AgCl  +  Na2SO,. 
2HC1    +2Ag  +  0=2AgCl  +  H20. 
CI    +  Ag        =  AgCl. 

Any  or  all  of  these  reactions  may  take  place  at  the  same  time.  The 
salt,  reacting  with  the  sulphates  of  iron  and  copper,  converts  those  metals 
into  their  higher  chlorides,  while  the  chlorine  and  hydrochloric  acid  are 
formed  at  the  same  time.  Both  chlorine  and  hydrochloric  acid,  at  the 
temperature  of  the  roasting  furnace,  react  readily  with  metallic  silver  or 
its  sulphate,  to  form  the  silver  chloride,  while  the  chlorides  of  iron  and 
copper,  in  chloridizing  the  silver,  may  pass  repeatedly  from  the  ferric 
and  cupric  condition  to  that  of  the  ferrous  and  cuprous: 

2FeCl2  +  2Cl  =2FeCl3. 
2FeCl3  +  2Ag  =  2FeCl2  +  2  AgCl. 
2CuCl  +2C1  =2CuCl2. 
2CUCI2  +  2Ag  =  Cu^Cl,  +  2  AgCl. 

The  ferric  chloride,  FeClj,  is  volatile  and  at  a  red  heat,  chloridizes 
the  silver  with  great  avidity.  The  ferrous  chloride  at  the  same  time  is 
resolved  into  ferric  oxide  and  ferric  chloride: 

3FeCl2  +  03  =  Fe203  +  2FeCl3. 

In  contact  with  aqueous  vapor,  and  the  fuel  gases,  at  a  red  heat,  the 
ferrous  chloride  may  be  converted  into  the  magnetic  oxide : 

3FeCl2  +  4H20=Fe30,  +  6HCl  +  2H. 

The  magnetic  oxide  may  be  again  reconverted  into  the  ferric  oxide, 
in  the  presence  of  salt  and  at  a  lower  temperature,  as  was  shown  con- 
clusively by  Stetefeldt'  who  succeeded  in  converting  an  ore  containing 
67.2  per  cent,  magnetite  to  1.4  per  cent,  after  4  1/2  hours'  roasting 
with  5  per  cent.  salt. 

Cupric  chloride  (CuClj)  is  easily  decomposed  at  a  red  heat  into  cuprous 

"Trans.  A.  I.  M.  E.,  1885-1886. 


chloride  (CujClj)  and  free  chlorine,  which  gives  free  chlorine  available 
for  the  chloridization  of  the  silver. 

Arsenic  and  antimony  form  chlorides,  which  are  easily  volatile 
and  which  may  be  decomposed  into  arsenous  and  antimonous  acids 
and  chlorine  and  hydrochloric  acid,  by  means  of  the  oxygen,  and  the 
vapor  from  the  burning  fuel.  These  chlorides,  however,  will  mostly 
escape  withput  decomposition.  If  the  temperature  is  low  and  the  salt 
has  not  been  added  until  the  arsenic  and  antimony  have  been  largely 
driven  off,  the  soluble  arsenates  and  antimonates,  in  the  roasted  ore, 
will  not  usually  be  present  in  sufficient  quantities  to  seriously  interfere 
with  the  extraction.  If  the  raw  ore  contains  arsenic  and  antimony  in 
large  amounts,  much  of  the  silver  may  be  converted  in  the  early  stages 
of  the  roasting,  into  arsenate  and  antimonate.  Ottokar  Hofmann  found 
in  roasting  arsenical  ore  that  53.8  per  cent,  of  the  silver  was  soluble  in 
sodium  hyposulphite,  probably  as  arsenate  of  silver,  before  the  salt  was 
added  to  the  ore. 

Arsenous  oxide  volatilizes  at  218°  C.  Chlorine,  with  the  aid  of  heat, 
decomposes  the  sulphide  of  antimony  completely,  forming  the  trichloride 
of  antimony  and  sulphur  dioxide.  The  trichloride  of  antimony  melts  at 
70°  C. 

Zinc. — Zinc  blende,  in  oxidizing  roasting,  is  converted  into  zinc 
oxide  and  zinc  sulphate,  while  sulphur  dioxide  escapes.  In  the  presence 
of  salt,  zinc  blende  remains  indifferent  and  does  not  decompose  salt,  at 
least  at  the  temperature  used  in  chloridizing  roasting.  Salt  does  not 
decompose  zinc  sulphate.  Zinc  oxide  may  be  converted  into  the  chloride 
at  a  red  heat.  By  the  action  of  chlorine  and  hydrochloric  acid  zinc 
chloride  is  formed,  which  is  very  volatile.  In  ore  which  has  been  given 
a  chloridizing  roast,  the  zinc  is  usually  found  as  the  oxide,  sulphate,  and 
chloride.  Zinc  oxide,  like  the  calcium  and  magnesium  oxides,  is  com- 
pletely soluble  in  acids,  so  that  when  an  acid  process  is  employed  to 
extract  copper  or  silver,  zinc  must  be  regarded  more  or  less  as  equivalent 
to  calcium  and  magnesium.  In  the  chlorination  of  gold  ores,  when  the 
chlorine  is  applied  direct  without  the  use  of  acids,  considerable  quanti- 
ties of  zinc  will  not  seriously  interfere  with  the  treatment.  Both  zinc 
oxide  and  zinc  sulphate  react  very  slowly  with  chlorine. 

Lead. — If  galena  is  subjected  to  chloridizing  roasting,  especially  in  the 
presence  of  sufficient  air,  most  of  the  lead  is  converted  into  sulphate, 
which  does  not  react  on  the  salt,  and  oxide,  which  may  be  converted 
into  the  chloride.  Both  lead  oxide  and  chloride  are  volatile,  while  the 
sulphate  remains  indifferent.  In  the  roasted  ore,  the  lead  will  be  in  the 
form  of  sulphate  and  chloride,  but  the  sulphate  will  predominate. 

Calcium  Carbonate. — Carbonate  of  lime,  when  roasted  with  metallic 
sulphides,  will  change  partly  into  calcium  sulphate  and  partly  into  the 
oxide  (lime).     The  calcium  sulphate  does  not  act  on  salt,  but  the  oxide 



decomposes  the  metal  sulphates  and  chlorides,  and  also,  to  some  extent, 
the  silver  chloride.  Calcium  oxide,  or  carbonate,  does  not  absorb 
chlorine  when  cold,  but  at  a  red  heat  combines  with  it  to  form  calcium 
chloride,  with  the  evolution  of  oxygen.  If  the  ore  contains  calcium 
carbonate  in  large  excess,  only  a  small  quantity  of  iron  and  copper  sul- 
phates will  be  formed,  to  decompose  the  sodium  chloride.  Most  of  the 
iron  and  copper  sulphides  in  the  ore  will  be  converted  directly  into  the 
oxides.  Since  the  sulphates  of  iron  or  copper  are  necessary  to  release  the 
chlorine  in  the  salt,  and  these  sulphates  are  not  formed  or  are  immediately 
appropriated  by  the  lime,  the  salt  will  not  react  to  release  chlorine  or  hy- 
drochloric acid,  which  are  the  most  active  elements  in  chloridizing  roast- 
ing of  copper,  gold,  and  silver  ores.  The  lime  itself  is  quite  indifferent  to 
silver  chloride  at  low  temperatures,  but  decomposes  it  energetically 
when  the  temperature  reaches  red  heat.  If  there  are  more  sulphides  in 
the  ore  than  are  necessary  to  convert  the  lime  into  sulphate  or  chloride, 
usually  a  good  chloridization  of  the  silver  and  copper  may  be  obtained. 
The  practical  effect  of  lime,  in  the  formation  of  silver  chloride,  in 
chloridizing  roasting,  is  clearly  shown  by  a  well  conceived  experiment  by 
Ottokar  Hofmann'  on  concentrates  containing  large  quantities  of  sul- 
phur, arsenic,  iron,  considerable  zinc,  some  lead  and  aluminum.  The 
object  of  the  experiment  was  to  ascertain  the  effect  of  varying  quantities 
of  calcium  carbonate,  in  the  formation  of  silver  chloride,  all  other  condi- 
tions remaining  the  same.  The  ore  was  roasted  one-half  hour  with  7 
per  cent.  salt. 


Per  cent,  of  con- 
centrates in 


Per  cent,  of  barren 

gangue  in  mixture 

mostly  CaCOg 


Value  of  mixture 
per  ton  02.  silver 


Value  of  leach 

tails  per  ton  oz. 


tion per  cent. 



The  deleterious  effect  of  the  lime  is  very  evident  from  these  results. 

Magnesium  usually  occurs  as  the  carbonate,  and  in  chloridizing 
roasting,  as  in  oxidizing  roasting,  has  about  the  same  effect  as  calcium. 
Magnesium  carbonate  is  decomposed  at  170°  C.  (338°  F.)  into  magnesium 
oxide.  If  there  are  sulphides  in  the  ore,  much  of  the  magnesium  will  be 
converted  into  the  sulphate.  The  sulphate  is  quite  infusible,  melting 
only  at  about  1 100°  C.  In  chloridizing  roasting  the  magnesium  combines 
with  the  chlorine  to  form  magnesium  chloride  (MgClj),  with  the  liberation 
of  oxygen.  Magnesium  chloride  fuses  at  a  red  heat,  708°  C.  (1300°  ¥.). 
Magnesium  chloride  is  more  positive  in  its  action  than  sodium  chloride. 

'  Min.  Ind.,  1896. 


Quartz.— Quartz  is  the  most  desirable  gangue  in  chloridizing,   as 
it  is  in  oxidizing  roasting.     Silica  is  indifferent  to  any  action  in  chloridiz-  ■ 
ing  roasting,  unless  perhaps,  it  promotes  the  formation  of  chlorides  and 
oxides  by  catalytic  action. 

Barium  sulphate,  which  occurs  quite  frequently  associated  with 
silver  ores,  remains  inert  during  chloridizing  roasting. 

Alumina  is  not  fused  by  heat  alone,  nor  is  it  decomposed  by  chlorine 
at  any  temperature. 

Sodium  sulphate,  so  abundantly  formed  during  chloridizing  roasting, 
may  be  considered  a  neutral  substance  in  any  of  the  hydrometallurgical 
processes.     It  is  usually  filtered  off  before  the  solvent  is  applied. 

Silver  sulphate  is  completely  decomposed  by  sodium  chloride,  at  the 
temperature  of  the  roasting  furnace. 

Percentage  of  Salt. — The  percentage  of  salt  used  at  various  mills 
differs  greatly,  depending  largely  on  the  character  of  the  ore,  principally 
the  gangue.  More  salt  has  sometimes  been  used  than  was  really  needed. 
Aaron  when  roasting  a  pyritic  ore  with  4. per  cent,  salt,  found  an  enor- 
mous loss  by  volatilization;  later  he  reduced  the  amount  of  salt  to  3  lb. 
per  ton  of  ore,  and  got  satisfactory  results.  Ordinarily,  the  amount  of 
salt  for  silver  ores  will  vary  between  1  and  5  per  cent.,  although  much 
greater  percentages  than  these  have  been  used.  Only  3  per  cent,  was 
used  at  Panimint,  California,  and  gave  a  chloridiz ation  of  95  per 
cent.  The  amount  of  salt  is  largely  proportional  to  the  amount  of 
lime  or  magnesia  in  the  ore.  An  excess  of  salt  does  not  improve  the 

If  copper,  instead  of  silver  ores  are  to  be  chloridized,  the  amount  of 
salt  required  will  be  larger.  If  the  copper  contained  in  the  ore  is  consider- 
able, the  amount  of  salt  will  be  roughly  proportional  to  the  copper.  From 
5  to  10  per  cent,  might  be  considered  fair  averages  for  ores  containing 
only  several  per  cent,  of  copper. 

The  minimum  amount  of  salt  that  may  be  used  for  any  ore  is  best 
determined  by  direct  experimenting.  First  determine  the  conditions  of 
time,  temperature,  and  fineness  of  the  ore,  which  will  give  the  highest 
satisfactory  chloridization  with  an  abundance  of  salt,  and  then  reduce  it 
in  successive  roasts  until  a  minimum  is  obtained  which  will  show  no  ap- 
preciable difference  as  compared  with  the  highest  chloridization  possible, 
with  an  abundance  of  salt. 

Time  of  Adding  Salt. — The  time  of  adding  salt  is  governed  almost 
entirely  by  the  composition  of  the  ore.  If  the  ore  is  low  in  sulphur,  the 
salt  may  be  added  before  the  ore  is  charged  into  the  furnace,  preferably 
before  it  is  crushed,  so  as  to  get  an  intimate  mixture  of  ore  and  salt. 
If  the  ore  contains  considerable  sulphur,  combined  with  iron  or  copper, 
the  ore  may  be  given  practically  a  full  oxidizing  roast  before  adding  the 
salt,  and  still  have  enough  sulphur  in  the  ore  to  chloridize  the  silver.     If 


copper  is  to  be  chloiidized,  the  ore  should  contain  at  least  as  much 
sulphur  as  copper  before  the  salt  is  added.  'If  the  ore  contains  consider- 
able zinc  and  lead  sulphides,  the  sulphur  combined  with  the  zinc  and  lead 
may  be  disregarded  for  the  purpose  of  chloridization,  and  the  ore  given 
an  oxidizing  roast  previous  to  adding  salt,  if  the  sulphur  combined  with 
the  iron  and  copper  is  large;  if  the  sulphur  so  combined  is  small,  the 
salt  is  best  added  at  once  to  the  raw  ore. 

If  the  ore  is  thoroughly  oxidized,  and  does  not  contain  sufficient  sul- 
phur, either  as  raw  ore,  or  after  a  thorough  oxidizing  roast,  the  salt  and 
pyrites,  both  finely  ground,  may  be  added  to  the  ore.  Ferrous  sulphate 
may  be  used  instead  of  pyritesj  but  is  much  more  expensive. 

If  the  salt  is  added  while  there  is  a  large  excess  of  sulphur  in  the  ore, 
it  will  largely  be  volatilized  without  doing  any  good.  If  the  salt  is  added 
at  the  proper  time,  the  chloridization  takes  place  very  rapidly. 

Heap  Chloridization. — Imperfectly  roasted  ore,  after  being  drawn  from 
the  furnace  and  placed  in  a  mass  on  the  cooling  floor,  or  in  a  pit,  will  gain 
in  chloridization,  largely  in  proportion  to  the  imperfectness  of  the  roast. 
On  very  poorly  roasted  ores  it  may  gain  as  much  as  50  and  75  per  cent. 
The  reactions  which  take  place  in  heap  chloridization  are  essentially  the 
same  as  those  which  take  place  in  the  furnace.  In  any  well  regulated 
mill,  the  ore  is  probably  never  so  poorly  roasted  but  that  all  the  iron  and 
copper  sulphides  are  decomposed. 

From  the  reactions  given  for  chloridizing  roasting,  it  is  evident  that 
air  is  not  essential  to  the  chloridization  after  the  sulphides  have  been 
converted  into  the  sulphates.  Small  quantities  of  air,  however,  permeate 
the  mass  and  promote  the  reactions. 

Ores  which  are  well  roasted  in  the  furnace,  and  which  is  the  only  safe 
way  to  roast,  do  not  show  any  increase  in  chloridization  in  the  heaps,  pit, 
or  cooling  floor. 

If  the  ore  does  not  contain  lime  in  considerable  quantity,  moistening 
the  hot  ore  adds  to  the  chloridization  of  the  silver  and  this  is  especially 
the  case  if  the  ore  contains  copper,  or  is  moistened  with  a  solution  of 
cupric  chloride.  If  the  ore  contains  appreciable  quantities  of  lime,  then 
instead  of  adding  to  the  chloridization,  there  is  likely  to  be  a  diminution. 
A  loss  of  chloridization  of  10  per  cent,  has  been  known  to  occur  in  this 

Composition  of  the  Roasted  Ore. — Ores  which  have  been  subjected  to 
chloridizing  roasting  contain  a  great  number  of  soluble  salts.  Of  these, 
sodium  sulphate,  resulting  from  the  decomposition  of  the  salt,  and  the 
undecomposed  sodium  chloride,  predominate.  Besides  these  there  may 
be  the  sulphates  of  manganese,  zinc,  copper,  iron,  aluminum,  and  mag- 
nesium; the  chlorides  of  the  same  metals  and  of  calcium  and  barium. 
The  barium  chloride  will  be  immediately  decomposed  on  solution,  and 
precipitated   as  insoluble  barium  sulphate.     Sodium  arsenate  is   also 



present  if  the  ore  contains  arsenic.  Salts  not  easily  soluble  in  water,  are 
cuprous  chloride,  lead  chloride,  calcium  sulphate,  sodium  antimonate, 
and  calcium  oxide.  Lead  chloride,  on  solution,  will  be  precipitated  as 
lead  sulphate.  Silver  chloride,  lead  sulphate,  and  antimonate*  are  al- 
most insoluble  in  water  but  are  soluble  in  solutions  of  other  chlorides. 
Cuprous  chloride,  calcium  sulphate  and  calcium  oxide  are  more  soluble 
in  a  chloride  solution  than  in  water.  If  the  ore  contains  large  quantities 
of  lime,  the  soluble  metals  may  be  precipitated  as  hydroxides. 

The  composition  of  Ontario  raw  ore,  and  of  the  ore  roasted  with  13 
per  cent,  salt  in  a  Stetefeldt  furnace,  from  analyses  made  by  Stetefeldt, 
is  given  by  Kustel'  as  follows: 











Manganese . 




Sulphur. . . . 


Antimony.  . 


Arsenic .... 




Alumina.  .  . 

13. U 




Bismuth.  .  . 


Cadmium.. . 




Magnesia. . . 


Roasted  ore;  shaft 


Roasted  ore;  Hue 


Zinc  chloride 

Copper  chloride 

Aluminum  chloride 

Sodium  chloride 

Traces  of  chlorides  of  other 

Aluminum  sulphate 

Lead  sulphate 

Sodium  sulphate 

Traces  of  sulphates  of  other 

Rest,  metallic  oxides  and 

Sulphur  in  undecomposed  sul- 




Aluminum  chloride 

Sodium  chloride 

Traces  of  chlorides  of  other  metals. 

Aluminum  sulphate 

Lead  sulphate 

Sodium  sulphate 

Copper  sulphate 

Zinc  sulphate 

Traces  of  sulphates  of  other 

Rest,  metallic  oxides  and  gangue 

Sulphur  in  undecomposed  sul- 




These  results  are  interesting  as  showing  the  condition  of  the  various 
constituents  of  the  ore,  after  chloridizing  roasting. 

Of  the  silver  contained  in  the  ore,  81.32  per  cent,  was  chloridized. 

Volatilization  of  the  Silver. — The  volatilization  of  the  silver,  in  chlorid- 
izing roasting,  is  largely  due  to  the  presence  of  other  chlorides  which 
are  more  volatile  than  the  silver  chloride.  The  volatilization  of  the 
silver  is  roughly  proportional  to  the  volatilization  of  the  base  metal 
chlorides,  or  to  the  loss  in  weight  the  ore  sustains.  Manganese  seems  to 
be  particularly  active  in  causing  loss  by  volatilization.  Cupric  and 
cuprous  chlorides,  both  of  which  volatilize  at  a  low  heat,  are  likely  to 
cause  a  heavy  loss  of  silver.  Arsenic  and  antimony  are  also  effective  in 
assisting  the  volatilization  of  the  silver  chloride.  A  high  temperature 
indirectly  causes  a  high  loss  of  silver  by  the  expulsion  of  the  volatile 

Much  of  the  loss  due  to  volatilization,  is  chargeable  to  the  manipula- 

'  Kustel,  "Roasting  of  Gold  and  Silver  Ores". 


tion  of  the  ore  in  the  furnace.  Any  condition  which  wUl  produce  the 
chloridization  of  the  silver,  if  carried  to  excess,  will  also  cause  its  volatil- 
ization. Silver  chloride,  under  the  conditions  of  roasting,  is  formed  at  a 
comparatively  low  temperature  by  the  chemical  reaction  between  the 
salt  and  sulphates.  A  scarcely  visible  red  heat  is  quite  sufficient  for 
these  reactions,  and  if  this  temperature  is  not  exceeded,  only  a  small  loss 
by  volatilization  will  occur.  If,  however,  the  temperature  is  elevated  to, 
say,  a  bright  red,  a  high  loss  of  silver  is  sure  to  take  place.  A  safe  rule  to 
follow,  is  to  keep  the  ore  at  the  lowest  possible  temperature  at  which  it 
will  give  off  visible  fumes.  It  is  best  to  maintain  a  deep  layer  of  ore,  and 
plenty  of  air.  A  small  charge  of  ore  spread  thinly  over  a  large  hearth 
area,  will  show  a  greater  loss  by  volatilization,  than  a  large  charge  with 
a  deep  bed  spread  over  the  same  area. 

The  stirring  of  the  ore  should  not  be  too  frequent,  but  this  is  not  an 
essential  if  the  temperature  is  not  too  high.  These  conditions  for  good 
chloridizing  roasting  are  contrary  to  those  desirable  in  the  best  oxidizing 
roasting,  where  the  ore  should  be  in  a  thin  layer  and  be  stirred  as  fre- 
quently as  possible.  Chloridizing  roasting  can  be  done  with  a  very  small 
loss  by  volatilization — frequently  only  an  inappreciable  loss — and  it  is 
very  probable  that  the  great  losses  recorded  are  due  entirely  to  improper 

In  the  chloridizing  roasting  of  any  ores,  at  an  exceedingly  low  temper- 
ature, a  difficulty  may  arise,  in  the  subsequent  chemical  treatment.  If 
roasted  at  too  low  a  temperature,  some  of  the  injurious  elements  may  not 
be  decomposed  sufficiently,  so  that  trouble  may  arise  in  the  consumption 
of  chemicals  when  the  solvent  is  applied  to  the  ore.  This,  however,  is  a 
matter  for  adjustment  for  each  particular  ore,  and  will  usually,  in  such 
cases,  resolve  itself  down  to  roasting  at  the  highest  temperature  the  ore 
will  stand  without  serious  loss  by  volatilization. 

Almost  any  ore,  likely  to  be  treated  by  a  solvent  process,  can  be 
effectively  chloridized,  but  the  essential  of  such  roasting  is  that  the  loss 
during  the  process  should  not  be  serious,  or  if  serious,  its  recovery  should 
be  carefully  considered.  With  care,  many  silver  ores  can  be  given  a 
chloridizing  roasting  with  not  much  greater  loss  of  silver  than  in  oxidizing 

That  there  is  sometimes  a  considerable  loss  of  silver  in  oxidizing  roast- 
ing is  pretty  well  established.  Plattner  in  his  "  Metallurgische  Rost- 
prozesse"  goes  very  minutely  into  the  loss  of  gold  and  silver  in  oxidizing 
roasting.  By  a  series  of  mufile  roasts  on  a  small  scale,  he  comes  to  the 
conclusion  that  while  there  is  no  loss  of  gold,  the  loss  of  silver  is  unavoid- 
able. From  numerous  tests,  varying  from  3/4  to  1  1/2  hours  he  records 
a  loss  of  from  0.5  to  18  per  cent,  of  the  silver.  He  concludes  that  the 
percentage  loss  of  silver  increases  with  the  temperature,  the  porosity  of 
the  charge  which  facilitates  the  supply  of  air  throughout  the  ore  mass,  the 



freedom  of  the  silver  from  combination  with  other  substances,  and  with 
the  time  of  the  roasting. 

In  order  to  verify  the  work  done  by  Plattner,  Christy'  cites  some 
experiments  in .  oxidizing  roasting  made  by  himself  and  others.  The 
material  used  in  the  experiments  were  concentrates  from  Nevada  City, 
California,  which  consisted  chiefly  of  pyrite,  with  small  amounts  of 
chalcopyrite  (0.05  to  1.5  per  cent.  Cu),  a  little  galena,  a  small  amount 
of  quartz,  traces  of  arsenic  and  antimony,  but  no  tellurium.  The  ore  was 
given  an  oxidizing  roast  of  from  1  1/2  to  8  1/2  hours;  in  the  early  stages 
at  incipient  dull  red,  and  finished  at  dull  red  to  full  red.  The  results 
are  tabulated  as  follows: 

Time  of 

Raw  ore,  ounces  per  ton 

Roasted  ore,  ounces  per  ton 

Percentage  loss  per  ton 

roasting,  hours 



Gold                 Silver 

Gold         1        Silver 

1  1/2 

2  1/2 

8  1/2 

4.. 58 






These  results  verify  the  conclusion  of  Plattner  and  others,  that  while  no 
loss  of  gold  occurs  in  oxidizing  roasting,  by  volatilization,  the  loss  of 
silver  may  be  considerable. 

Butters^  found  in  roasting  a  hard  white  quartz,  intimately  mixed 
with  about  7  per  cent,  calcite  and  a  very  little  pyrite,  assaying  5.55  oz. 
silver  and  0.65  oz.  of  gold,  per  ton,  that  there  was  a  loss  by  volatilization 
in  oxidizing  roasting,  of  2  to  9  per  cent,  of  the  silver,  but  none  of  the 

It  is  possible  that  losses  of  silver,  which  have  been  attributed  to 
chloridizing  roasting  may  have  been  partly  due  to  the  loss  in  oxidizing 
roasting,  and  especially  if  some  of  the  silver  in  the  ore  is  in  the  form  of 

Volatilization  of  the  Gold. — It  is  pretty  well  established  both  by  care- 
fully conducted  experiments  and  by  the  experience  of  practical  metal- 
lurgists, that  no  loss  of  gold  takes  place  either  in  oxide  or  sulphide  ores, 
in  oxidizing  roasting.  There  seems  to  be  some  doubt  in  the  case  of  tellur- 
ides,  but  the  experience  with  Cripple  Creek  ores,  containing  tellurium,  of 
which  hundreds  of  tons  are  roasted  daily,  is,  that  no  appreciable  loss,  if 
any  at  all,  occurs  by  volatilization.  Kustel  records  a  loss  of  20  per  cent, 
of  the  gold  during  the  oxidizing  roasting  of  certain  telluride  ores  of  gold 
and  silver,  and  states  that  this  is  not  a  mechanical  but  a  volatilization 
loss.     There  can  be  no  doubt  about  the  gold,  combined  with  tellurium, 

>  Trans.  A.  I.  M.  E.,  88-89. 
2  Trans.  A.  I.  M.  E.,  88-89. 


volatilizing  at  elevated  temperatures,  but  whether  any  volatilization 
takes  place  at  the  low  temperatures  and  under  the  practical  conditions 
of  roasting,  seems  very  doubtful. 

Tellurides,  even  in  small  quantities,  are  extremely  sensitive  to  chlorine 
at  almost  any  temperature,  at  which  salt  is  decomposed.  Experience 
with  Cripple  Creek  ores,  in  large  100-ton  furnaces,  showed  appreciable 
loss  of  gold  when  only  a  very  small  amount  of  salt — from  1/2  to  2  per 
cent. — was  added  during  the  roasting.  A  loss  was  shown  even  when  the 
salt  was  added  to  the  hot  ore  dropping  on  the  cooling  hearth. 

Many  ores  are  known  to  contain  chlorine,  frequently  as  chloride  of 
silver  or  chloride  of  lead.  That  ore  containing  a  part  of  its  silver  as 
chloride,  if  given  an  oxidizing  roast,  will  volatilize  small  amounts  of  both 
gold  and  silver,  was  proved  conclusively  by  the  author  in  exhaustive 
tests  on  Tonapah  ore.  It  is  probable  when  gold  losses  occur  in  any  ore 
in  oxidizing  roasting,  and  especially  in  the  tellurides,  it  may  be  due  to 
small  quantities  of  chlorine. 

Prof.  Christy'  made  some  interesting  experiments  on  the  volatiliza- 
tion of  gold  in  the  chloridizing  roasting  of  pyritic  ores.  As  the  result  of 
a  large  number  of  experiments  he  comes  to  the  following  conclusions: 

At  100°  C.  (212°  F.)  the  volatility  of  the  gold  in  an  atmosphere  of 
chlorine,  is  almost  zero;  that  the  loss  begins,  above  this  temperature,  to 
rapidly  increase  to  a  maximum  at  a  temperature  of  about  250°  C.  (482° 
F.);  that  it  rapidly  diminishes  to  a  temperature  somewhere  below  red 
heat;  that  it  again  increases,  but  more  slowly,  to  another  maximum,  at 
a  temperature  above  a  melting  heat,  and  that  this  increase  is  apparently 
continuous  between  a  red  heat  and  a  white  heat.  The  ratio  of  losses  at 
various  temperatures  is  also  instructive;  at  incipient  redness  the  standard 
loss  is  already  0.05  per  cent.;  at  a  cherry  red  it  is  five  to  seven  times  as 
great  as  at  incipient  redness;  at  incipient  yellow  it  is  more  than  eight 
times  what  it  is  at  incipient  red;  while  at  melting  heat  it  is  nearly  thirty 
times  as  great. 

Crosly^  found  with  a  certain  California  pyritic  ore,  assaying  about 
$110.00  in  gold,  and  $40.00  in  silver,  that'  an  oxidizing  roast  showed  no 
appreciable  loss,  but  when  the  salt  was  added,  losses  appeared  rapidly. 
Thus,  according  to  his  tests,  with  3  per  cent,  salt  the  gold  loss  was  30 
per  cent,  and  the  silver  loss  50  per  cent,  of  the  assay  value.  He  at- 
tributed the  loss  to  the  presence  of  tellurides,  which  he  supposed  were 

Aaron^  found  a  large  loss  in  roasting  a  simple  pyrite  in  a  3-hearth 
reverberatory  furnace,  with  1  to  2  per  cent,  of  salt,  which  was  added 
on  account  of  the  silver.     He  then  made  two  tests  on  a  small  scale; 

'  Trans.  A.  I.  M.  E.,  1885. 
'  Trans.  A.  I.  M.  E.,  1888. 
=  "Leaching  Gold  and  Silver  Ores,"  1881. 


one  with  4  per  cent,  salt,  the  other  without  any  salt,  and  purposely 
pushed  the  roasting  to  an  extreme  as  to  time  and  temperature,  and 
found  on  assaying  that  the  salted  ore  contained  less  than  half  as 
much  gold  as  the  unsalted  one.  He  also  found  that  the  ore,  in 
roasting,  sustained  a  loss  of  18  per  cent,  in  weight,  and  consequently 
should  have  assayed  18  per  cent,  more  than  the  raw  ore,  which  was  not 
the  case.  By  modifying  the  roasting,  so  as  not  to  add  the  salt  until  the 
dead  roasting  of  the  ore  was  finished,  not  only  did  the  roasted  ore  assay 
20  per  cent,  more  than  the  raw  ore,  but  the  yield  overran  the  guarantee, 
while  the  tailings,  nevertheless,  contained  considerably  more  gold  than 
before.  He  afterward  found  that  a  very  small  quantity  of  salt — not  more 
than  3  lb.  per  ton  of  ore — might  be  mixed  with  the  raw  ore  without 
detriment  to  the  gold  and  with  decided  advantage  to  the  extraction  of 
the  silver. 

The  principal  object  of  roasting  gold  ores,  containing  silver  or  copper, 
with  salt,  is  to  chloridize  the  small  amounts  of  silver  and  copper,  and  in 
some  cases  to  neutralize  substances  in  the  ore,  which  might  be  injurious 
to  the  solvent.  By  a  partial  chloridizing  roast,  or  even  with  an  oxidizing 
roast,  it  is  practicable  to  get  a  high  extraction  of  both  the  gold  and  silver 
by  either  the  cyanide  or  chlorination  processes.  If  it  is  simply  a  matter 
of  neutralizing  injurious  substances  in  the  ore,  this  can  be  done  in 
chloridizing  roasting  by  not  pushing  the  operation  to  the  limit,  and  if  not 
carried  beyond  the  point  required  to  satisfy  the  base  elements,  no 
appreciable  volatilization  of  either  gold  or  silver  will  occur. 

Chloridization  of  Copper  Ores. — According  to  Von  Kothny^  by  roast- 
ing copper  sulphide  mixed  with  iron  oxide  and  sodium  chloride  prac- 
tically all  the  sulphur  goes  into  sulphate  and  about  half  the  copper 
is  transformed  into  chloride.  Anhydrous  cupric  chloride  mixed  with 
sodium  chloride  and  heated  in  a  current  of  air  to  250°  C.  gives  off 
chlorine.  The  decomposition  of  copper  sulphate  by  sodium  chloride 
begins  at  280°  C.  Ferric  chloride  converts  copper  oxide  into  chloride 
very  rapidly  at  temperatures  from  500  to  600°  C.  The  formation 
of  copper  sulphate  by  roasting  with  copper  oxide  in  the  presence  of 
sodium  chloride  plays  no  part.  At  temperatures  of  300  to  600°  C.  ferric 
sulphate  converts  copper  oxide  slowly  into  sulphate.  Chlorine  is  with- 
out direct  action  on  cuprous  sulphide.  The  reactions  involved  in  the 
Hargreaves  process  by  which  hydrochloric  acid  is  formed  plays  no  part 
in  converting  cupric  oxide  into  chloride.  Von  Kothny  concludes  that 
the  mechanism  of  the  chloridizing  of  pyrite  cinder  containing  a  small 
amount  of  copper  and  sulphur,  is  as  follows :  The  copper  is  present  largely 
as  sulphide,  which  by  an  oxidizing  roasting  is  converted  into  sulphate  and 
oxide.  Sodium  chloride  acts  directly  on  the  sulphate  and  ferric  chloride 
on  the  oxide.  To  insure  chloridizing  of  such  material  it  must  be  finely 
•  Metallurgie,Ju\j  8,  1911. 


ground;  a  large  amount  of  air  must  be  admitted  in  the  oxidizing  roasting 
period  and  stirring  must  be  resorted  to  to  insure  contact  with  oxygen; 
sufficient  pyrite  must  be  present  to  furnish  the  required  amount  of  ferric 
chloride  and  is  best  added  in  a  weathered  form;  for  4  per  cent,  copper 
content  at  least  7.5  per  cent,  salt  must  be  added;  the  process  should  be 
carried^  out  at  temperatures  between  500  and  600°  C.  For  a  full  discussion 
of  chloridizing  roasting  of  copper  ores  see  "  Longmaid-Henderson  process, " 
Part  II,  page  246. 

Principal  Factors  in  the  Loss  of  Silver,  Gold,  and  Copper  by  Volatiliza- 
tion.— -The  principal  factors,  controling  the  loss  of  silver,  gold  and  copper 
by  volatilization,  in  chloridizing  roasting,  have  been  well  established  both 
by  practice  and  careful  experiments.  These,  in  the  order  of  their  im- 
portance, are: 

1.  Temperature. 

2.  Time. 

3.  Amount  of  air,  or  surface  exposed. 

The  amount  of  salt  has  some  influence  on  the  volatilization,  but  it  is 
supposed  that  the  amount  of  salt  used  is  the  least  that  will  give  satisfac- 
tory results,  and  once  determined,  becomes  constant. 

The  presence  of  volatile  substances,  such  as  arsenic,  antimony, 
selenium,  tellurium,  and  the  chlorides  of  copper  and  iron,  also  affect  the 
volatilization.  Gold  is  particularly  sensitive  to  tellurium  in  chloridizing 
roasting.  But  as  these  are  constituents  of  the  ore,  they  cannot  be  con- 
sidered as  variable,  or  controllable  factors,  except  in  so  far  as  preliminary 
oxidizing  roasting  may  eliminate  them. 

Temperature  is  the  all  important  factor  in  chloridizing  roasting.  Any 
ore  chloridized  at  an  excessive  heat  will  volatilize  much  of  the  metals, 
irrespective  of  any  considerable  time,  or  in  any  atmosphere  attainable  in 
a  roasting  furnace.  If  the  temperature  is  kept  at  the  lowest  possible 
point  at  which  the  metals  can  be  chloridized,  then  the  time  of  roasting 
and  the  amount  of  oxygen  in  the  furnace  atmosphere  i;i  immaterial.  By 
merely  changing  the  temperature,  from  10  to  80  per  cent,  of  the  metals 
may  be  volatilized  in  a  short  time;  or  only  a  few  per  cent,  may  be  vola- 
tilized after  several  hours  roasting,  all  other  conditions  remaining  the 

Russel,  experimenting  with  Ontario  ores,  found  a  volatilization  of  8.3 
per  cent,  of  the  silver  at  a  dark  red  heat,  and  of  17.6  per  cent,  at  a  cherry 
red.  Ottokar  Hofmann'  found  in  roasting  calcareous  ores  containing 
large  quantities  of  zinc  and  arsenic,  that  the  ore  lost  3.5  per  cent,  of 
its  weight  and  1.8  per  cent,  of  its  silver  was  volatilized  when  roasted 
at  a  low  temperature;  the  same  ore  roasted  at  a  high  temperature 
with  insufficient  air,  lost  7  to  13  per  cent,  of  its  weight,  and  15  to  25 

'  Min.  Ind.,  1896. 


per .  cent,  or  more  of  the  silver.  He  also  found'  on  an  ore  consisting 
essentially  of  25  per  cent,  zinc,  12  per  cent,  lead,  21  per  cent,  sulphur, 
7  per  cent,  iron,  and  10  per  cent,  calcium  carbonate,  that  the  loss  by 
volatilization  varied  from  1.7  to  15  per  cent.  The  least  increase  of  tem- 
perature above  a  dull  red,  caused  a  heavy  loss,  even  if  the  increase 
lasted  for  only  a  short  time.  The  average  of  31  days  roasting  at  a  high 
(almost  white)  heat  was: 

Chloridization  of  the  silver,  72.7    per  cent. 
Loss  by  volatilization,  17.9    per  cent. 

Roasted  at  a  low  heat  (not  above  a  dull  red) : 

Chloridization  of  the  silver,  81.5    per  cent. 
Loss  by  volatilization,  1.2    per  cent. 

The  chloridization  in  favor  of  the  lower  heat  was  8.8  per  cent,  and  a 
decrease  of  loss  by  volatilization  of  16.7  per  cent. 

Time. — The  volatilization  of  the  silver,  gold,  and  copper,  in  chloridiz- 
ing  roasting,  is  approximately  proportional  to  the  time  of  roasting,  other 
conditions  remaining  the  same.  If  in  chloridizing  roasting,  an  ore  will 
lose,  say  1  per  cent  in  the  first  hour  after  the  salt  is  added,  it  will  lose 
approximately  5  per  cent,  after  five  hours  roasting,  if  the  conditions 
remain  the  same. 

Air  or  Oxygen. — Time  and  temperature  remaining  the  same,  the  vol- 
atilization will  be  approximately  proportional  to  the  amount  of  air 
supplied  to  the  ore.  If  a  ton  of  ore  is  roasted  on  a  hearth  area  of  100  sq.  ft. , 
shows  a  volatilization  of  say,  1  per  cent,  per  hour,  it  is  likely  to  show  2 
per  cent,  per  hour  if  spread  over  a  hearth  area  of  200  sq.  ft. 

Experiments  as  Compared  with  Practice. — Almost  all  the  chloridizing 
roasts  made  in  preliminary  tests,  in  a  muffle,  will  show  a  higher  loss  by 
volatilization  than  will  subsequently  be  found  in  practice.  There  is  no 
appreciable  loss  in  heap  chloridization,  when  improperly  roasted  ore  is 
withdrawn  from  the  furnace  and  the  chloridizing  allowed  to  proceed  on 
the  cooling  floor.  Neither  does  any  appreciable  loss  occur  when  the 
damper  of  the  furnace  is  closed,  so  that  the  furnace  has  no  draft  and  no 
fresh  air  supply. 

Relation  of  Sulphur  to  the  Chloridization  of  Silver  and  Gold. — Chlorid- 
ization of  the  silver,  in  chloridizing  roasting,  may  take  place  very  rapidly 
under  proper  conditions.  If  the^  ore  contains  an  excess  of  sulphur, 
chloridization  will  not  take  place  to  any  appreciable  extent,  until  some 
of  the  sulphur  has  been  eliminated,  even  if  there  are  sulphates  present. 
This  may  be  due  to  the  reducing  action  of  sulphur  dioxide,  or  other  reduc- 
ing gases,  which  are  likely  to  occur  in  abundance  in  the  early  stages  of 
the  roasting. 

'  Engineering  and  Mining  Journal,  1888-89. 



In  order  to  determine  the  relation  of  the  chloridization  of  the  silver 
and  gold  to  the  sulphur,  and  the  progress  of  chloridization  during  the 
roasting,  the  following  interesting  results  were  obtained  by  the  author 
on  Tonapah  concentrates,  which  consisted  largely  of  silica  and  iron 
pyrites,  about  3  per  cent,  lead,  and  small  quantities  of  zinc,  copper, 
manganese,  and  antimony.  The  raw  ore  had  16.35  per  cent,  sulphur,  and 
assayed  615.0  oz.  silver,  and  6.50  oz.  gold  per  ton.  The  roasting  was  done 
in  a  furnace  having  a  hearth  area  of  100  sq.  ft.  The  concentrates  were 
first  given  an  oxidizing  roast  for  two  hours,  after  which  10  per  cent, 
salt  was  added  and  samples  taken  every  hour. 




Sulphur,  per  cent. 

Roasted  ore,       ^        Hypo  tails, 
value  ounces      !      value  ounces 

per  cent. 

Total    I  Soluble  'insoluble'    Silver        Gold     ;    Silver         Gold         Silver        Gold 

0  hours 1  16.35 

1  hour 9.45 

2  hours i  7.10 

3  hours j  6.65 

4  hours I  6.45 

5  hours 6.25 

8  hours 6.15 



















It  will  be  noticed  that  no  chloridization  took  place  the  first  2  hours  of 
chloridizing  roasting,  notwithstanding  that  there  was  from  3.0  to  4.65 
per  cent,  soluble  sulphur  in  the  ore  at  that  time.  The  same  results  in 
chloridization  would  doubtless  have  been  obtained  if  the  salt  had  been 
added  three  hours  later  than  it  was,  or  after  the  ore  had  been  given  an 
oxidizing  roast  for  5  hours. 

The  high  value  of  the  hypo  tails  after  1  and  2  hours  chloridizing 
roasting,  is  due  to  the  fact  that  there  was  no  silver  chloride  formed,  and 
in  leaching  with  the  hypo  the  soluble  matter  was  removed,  thereby 
somewhat  concentrating  the  value  of  the  ore. 

The  high  soluble  sulphur  in  the  roasted  ore  is  mostly  due  to  the 
sodium  sulphate  formed  by  the  roasting.  Some  of  the  insoluble  sulphur 
may  have  been  in  the  form  of  lead  or  calcium  sulphate. 

Determination  of  Loss  by  Volatilization. — In  order  to  roast  skillfully 
it  is  of  great  importance  to  frequently  ascertain  the  loss  by  volatilization, 
but  to  do  this  it  is  necessary  to  know  the  loss  of  weight  the  ore  sustains. 
In  practical  handling  of  the  ore  this  is  difficult  and  inconvenient.  Hof- 
mann*  gives  the  following  method,  which  can  be  performed  in  an  assay 
office  in  a  few  hours : 

"Ten  grams  of  the  raw  pulp,  containing  the  same  percentage  of  salt 
as  the  ore  in  the  furnace,  is  placed  in  a  roasting  dish  and  roasted  in  the 

'  "Hydrometallurgy  of  Silver,"  Page  22. 


muffle  for  half  an  hour  or  an  hour;  then  the  sample  is  removed  from  the 
muffle,  allowed  to  cool,  weighed,  returned  to  the  muffle,  roasted  again  for 
half  an  hour,  and  then  weighed  again.  This  is  repeated  until  two  weigh- 
ings are  alike,  or  until  in  the  last  half  hour  the  ore  does  not  lose  more  than 
2  or  3  mg.,  then  the  difference  between  the  original  weight  and  that  of 
the  last  weighing,  expressed  in  percentage,  gives  the  highest  possible 
loss  the  raw  ore  can  suffer. 

Ten  grams  of  a  sample  of  roasted  ore,  corresponding  with  the  sample 
of  raw  pulp,  is  placed  in  a  roasting  dish,  and  also  roasted  in  the  muffle 
until  two  weighings  agree,  or  the  difference  between  two  consecutive 
weighings  is  not  more  than  2  or  3  mg.  The  difference  between  the  first 
weighing  (10  grm.)  and  the  last,  expressed  in  percentage,  gives  the  weight 
which  the  roasted  ore  is  still  capable  of  losing  if  subjected  to  prolonged 
roasting.  If  we  deduct,  therefore,  the  capable  loss  from  the  highest 
possible  loss,  we  obtain  in  percentage  the  loss  in  weight  the  ore  has 
suffered  during  roasting  in  the  furnace  by  volatilization." 

Chloridization  Determination. — To  determine  the  amount  of  silver 
and  gold  chloridized,  it  will  usually  be  sufficient  for  practical  purposes  to 
take  several  ounces  of  an  average  sample  of  the  ore,  and  treat  it  thor- 
oughly with  a  solution  of  sodium  hyposulphite.  It  will  be  found  most 
satisfactory  to  put  the  ore  and  hypo  in  a  beaker  for  several  hours  at  least, 
stirring  it  occasionally,  and  then  thoroughly  filter  and  wash  the  ore  in  a 
funnel.  It  is  then  dried,  bucked,  and  assayed.  If  the  ore  contains  large 
quantities  of  soluble  salts,  the  sample  should  be  weighed  before  and  after 
the  hypo  treatment,  and  the  difference  allowed  for  in  the  results  of  the 

Chloridization  determinations  are  sometimes  made  by  taking  an 
assay  ton,  or  less,  of  the  ore  treating  it  with  hypo,  and  assaying  the 
residue.  While  this  rectifies  any  error  of  soluble  salts,  it  introduces  a 
more  or  less  uncertain  element  in  the  assaying. 



Color  Names  of  Temperatures. — The  temperatures  corresponding  to 
different  colors  have  been  determined  quite  accurately  by  White  and 
Taylor,  by  Howe,  by  Janivier,  and  by  Pouillet.  The  difficulty  in  deter- 
mining a  certain  temperature,  by  its  corresponding  color,  lies  in  the 
personal  equation  of  the  observer  and  the  time  and  conditions  of  observa- 
tion. Much  depends  on  the  susceptibility  of  the  retina  of  the  observer 
to  light  as  well  as  the  degree  of  illumination  under  which  the  observation 
is  made.  A  furnace  looks  very  much  hotter  at  night  than  at  day,  and 
hotter  in  a  dark  room  than  in  a  bright  one.  The  most  experienced 
roasterman  is  unable  to  compensate  fully  for  these  factors,  nevertheless, 
the  information  given  by  these  color  temperatures  is  often  convenient. 

White  aad  Taylor 

Name  of  color 

Dark  red,  blood  red,  low  red. 

Dark  cherry  rod 

Full  cherry  red 

Light    cherry,  bright  cherry, 
bright  red. 


Light  orange 


Light  yellow 












1,176  J 



Name  of  color 

C.  F. 

Lowest  red  -visible  in  the  dark. 
Lowest  red  visible  in  daylight. 

Dull  red 

Full  cherry  . 
Light  red. . . 


FuU  yellow. . . 
Light  yellow. 




(£.  and  M.  J.,  Jan.  20,  1900.) 





Name  of  color 




Name  of  color 



Incipient  red 

Dull  red 

Incipient  cherry  red 

Cherry  red 

Clear  cherry  rod. . . . 
Deep  orange 

Clear  orange 


Bright  white 

Dazzling  white 







Very  dull  red .... 

Dull  red 

Bright  red. ...'... 

Cherry  red 

Bright  cherry  red 
Very  deep  orange 
Deep  orange  red . 

Orange  red 


Brilliant  white. . . 
Dazzling  white. . . 
Blue  white 














(E.  and  M.  J.,  July  20,  1905.) 

Pyrometric  Determinations. — The  only  way  of  accurately  determining 
the  temperatures  in  various  parts  of  a  roasting  furnace,  under  all  condi- 
tions, is  by  the  use  of  reliable  pyrometers,  and  every  roasting  plant  should 
be  equipped  with  at  least  one  of  these  instruments. 

Attempts  have  frequently  been  made  to  get  uniformity  in  the  quality 
of  ore,  roasted  for  treatment  by  the  chemical  processes,  by  establishing  a 
system  of  absolute  temperatures  in  certain  parts  of  the  furnace,  and  so 
firing  as  to  keep  those  temperatures  constant.  On  theoretical  grounds 
this  appears  quite  feasible.  The  difficulty  lies  in  assuming  that  the  ore 
fed  into  the  furnace  is  of  uniform  quality,  and  that  the  other  essential 
factors,  such  as  air  supply,  always  remain  the  same.  Ore  which  is  well 
bedded,  and  containing  about  2.5  per  cent,  sulphur,  may  vary  as  much  as 
0.35  per  cent,  to  0.50  per  cent.,  in  24  hours.  The  conditions  which 
would  be  ideal  for  ore  having  2.25  per  cent,  sulphur  would  be  far  from 
ideal  for  ore  having  2.50  or  2.75  per  cent. 

Much  also  depends  on  the  physical  and  chemical  composition  of  the 
ore.  The  condition  of  temperature  which  would  give  the  best  results  for 
partly  oxidized  ore  would  not  give  satisfactory  results  with  ore  containing 
the  same  amount  of  sulphur,  from  the  deeper  workings  of  the  mine,  in 
which  no  oxidation  had  taken  place,  assuming  of  course,  that  the  amount 
of  ore  roasted  remains  the  same.  When  ore  contains  an  excessive  amount 
of  dust,  it  cannot  be  roasted  at  the  same  rate  and  at  the  same  temperature 
as  ore  which  contains  only  the  normal  quantity,  and  the  dust  is  likely  to 
vary,  especially  when  the  supply  bins  get  low. 

The  temperature  of  a  roasting  furnace  appears  to  be  very  much 
hotter  at  night  than  during  the  day.     Inexperienced  roastermen  are 

PYROMETRY   '  81 

frequently  misled  by  this,  and  even  experienced  men  cannot  judge 
accurately  within  the  desired  limits.  The  tendency  in  the  daytime, 
especially  in  well  lighted  buildings,  is  to  get  the  temperature  too  high, 
and  at  night  to  get  it  too  low.  In  such  cases  pyrometers  are  of 
great  service  in  establishing  temperatures.  They  are  also  of  great 
service  in  determining  the  temperature  beyond  which  it  is  unsafe  to  roast. 
Experience  and  skill  in  the  appearance  of  the  ore  as  it  progresses  through 
the  furnace,  and  its  appearance  after  roasting,  however,  are  the  best  gen- 
eral guides  to  obtaining  uniform  results.  It  is  questionable  whether, 
even  with  a  perfect  system  of  pyrometry,  the  experience  and  skill  of  the 
operator  will  not  always  remain  the  dominant  factor. 

It  is  a  curious  fact  that  when  furnaces  are  overheated,  the  amount 
of  sulphur  in  the  roasted  ore  is  abnormally  high.  One  of  the  dangers  in 
employing  new  roastermen  is,  that  in  their  anxiety  to  get  a  good  roast, 
they  invariably  fire  at  too  high  a  temperature,  with  the  result,  that  the 
roasted  ore  contains  an  unusually  large  amount  of  sulphur,  partly  fused, 
and  is  in  the  worst  possible  condition  for  treatment  by  a  solvent  process. 
Overheated  partially  roasted  ore  is  also  likely  to  run  more  or  less  like  a 
liquid  and  in  this  way  emerge  insufficiently  roasted.  Pyrometers,  in 
such  cases,  are  invaluable  as  a  warning  to  the  roasterman  when  the  safe 
limit  of  temperature  is  being  exceeded.  Frequently  pyrometric  deter- 
minations are  essential  to  intelligent  work,  but  they  must  he  supple- 
mented by  experience  and  skill,  and  not  dominate  them. 

Of  the  pyrometers  in  general  use,  those  of  the  Le  Chatelier  type  will 
be  found  most  satisfactory  in  roasting  work.  If  the  thermo-electric 
couple  is  protected,  it  may  be  inserted  into  the  furnace  and  kept  there  a 
very  long  time — in  fact  more  or  less  permanently — without  appreciable 
injury.  The  limit  of  temperature  at  which  it  is  safe  to  use  these  pyrom- 
eters is  a  little  below  the  melting  point  of  platinum,  which  is  about  3250° 
F.,  although  readings  above  3000°  F.  cannot  be  relied  upon  as  perfectly 
accurate.  In  roasting  work  these  temperatures  are  never  approached. 
It  is  rarely  that  1700°  F.  (927°  C.)  is  exceeded.  From  1400°  F.  to  1600° 
F.  (760  to  871°  C.)  is  the  usual  range  in  the  hottest  part  of  the  furnace, 
for  the  various  ores.  Frequently  ores  are  encountered  which  give  the 
best  results  at  as  low  a  temperature  as  1000°  F.  (528°  C).  It  will  be 
seen,  therefore,  that  the  only  danger  to  the  thermo-electric  couple  of  the 
pyrometer  is  from  the  furnace  gases.  Even  this  danger  is  remote  in 
any  case,  and  is  entirely  obviated  if  the  thermo-electric  couple  is 

Several  pyrometers  inserted  at  various  points  of  a  large  roasting  fur- 
nace will  give  invaluable  information  as  to  the  limits  of  temperature, 
which  for  any  particular  ore,  will  give  the  best  extraction.  Once  these 
extreme  limits  have  been  determined,  it  is  an  easy  matter  to  fire  the 
furnace  so  that  they  shall  not  be  exceeded. 


While  the  determination  of  the  absolute  temperatures  in  roasting  is 
not  essential,  nevertheless  it  is  highly  desirable.  One  of  the  gratifying 
features  of  the  Le  Chatelier  type  of  pyrometer  is,  that  absolute  tempera- 
tures may  be  determined  with  greater  facility  and  accuracy  than  the 
relative  temperatures  may  be  determined  by  other  means,  and  absolute 
temperatures  are  always  reliable  for  comparison. 

The  elements  of  a  Le  Chatelier  pyrometer  consist 

1.  Of  a  thermo-electric  couple,  which  generates,  when  heated,  a 
slight  electric  current,  which  is  proportional  to  the  heat  applied. 

2.  A  galvanometer  so  arranged  that  the  deflection  of  the  needle,  due 
to  the  current,  indicates  the  temperature  on  the  scale  of  the  galvanometer. 

3.  Flexible  wires  connecting  the  thermo-electric  couple  with  the 

In  using  the  pyrometer,  the  thermo-electric  couple  may  be  inserted 
directly  into  the  furnace  at  the  points  where  the  temperature  is  desired, 
and  the  reading  taken.  Such  a  proceeding  is  awkward  and  troublesome. 
When  only  one  galvanometer  is  used,  the  most  satisfactory  arrangement  is 
to  permanently  insert  the  thermo-electric  couples  at  the  various  points 
of  the  furnace,  as  desired,  and  by  small  switches  and  the  wires  connecting 
the  couples  with  the  galvanometer,  the  temperatures  of  the  furnace  at 
the  different  po'nts  may  be  quickly  determined.  In  the  same  way,  the 
temperatures  of  different  furnaces  may  be  readily  ascertained.  The 
galvanometer  should  be  located  at  a  convenient  point,  away  from  the 
dust  and  fumes  of  the  furnace  room.  The  thermo-electric  couples  may  be 
inserted  into  the  furnace  through  the  arch,  but  care  must  be  taken  not  to 
project  them  down  far  enough  to  be  injured  by  the  rabbles,  although  it 
is  desirable  to  get  them  as  close  to  the  ore  as  possible.  The  thermo- 
couple, where  it  is  intended  to  remain  permanently  in  the  furnace,  should 
be  protected  by  porcelain  tubes. 

If  a  continuous  record  of  the  temperature  at  any  one  point  is  desired, 
it  is  best  to  use  a  recording  pyrometer.     This  consists  essentially  of: 

1.  A  recorder,  which  is  composed  of  a  galvanometer  and  a  clock 
arrangement,  so  that  a  pencil  indicates  the  temperature  and  time  on  a 
moving  chart. 

2.  The  thermo-electric  couple,  the  fire  end  of  which  is  inserted  into 
the  space,  where  the  temperature  is  to  be  measured. 

3.  Flexible   wires    connecting  the   recorder   with  the  thermo-couple. 
With  this  apparatus,  a  continuous,  automatic,  and  permanent  record 

of  temperature  and  time  may  be  made,  which  will  give  an  accurate  idea 
of  the  firing  of  the  roasterman  during  the  entire  shift.  A  comparison  of 
charts,  will  quickly  establish  the  best  temperature  at  which  it  is  desirable 
to  roast  any  ore,  and  locate  the  responsibility  of  any  defects  in  the  ore  due 
to  the  temperature  in  roasting. 


Roasting  furnace  design  as  applied  to  roasting  ores  preparatory  to 
treatment  by  the  hydrometallurgical  processes,  is  rapidly  resolving 
itself  down  to  the  various  types  of  mechanical  reverberatories.  Hand 
reverberatories  are  still  in  use  in  small  reduction  works,  but  even  for 
small  output  they  are  rapidly  being  displaced  by  the  more  efficient  me- 
chanical roasters.  Labor,  in  hand  roasting,  has  been  the  most  impor- 
tant factor  in  the  cost  of  operation,  especially  in  mining  districts,  where 
labor  is  from  $2.50  to  $4.00  a  day.  Besides,  the  quality  of  labor  is  an 
exceedingly  variable  factor,  and  as  the  quality  of  a  roast  depends  much 
on  the  efficiency,  conscientiousness,  and  skill  of  the  workmen,  it  is  a 
disturbing  element  in  any  metallurgical  plant  where  hand  roasting 
is  used. 

The  tendency,  therefore,  has  been  to  eliminate  the  labor  item,  and  the 
personal  factor,  by  the  substitution  of  mechanical  for.  hand  furnaces.  A 
mechanical  furnace,  once  set  in  motion  under  the  conditions  which  liy 
experiment  have  been  found  to  give  the  best  results,  will  alwaj-s  give 
practically  the  same  results  if  the  same  conditions  are  maintained,  and 
it  is  possible  to  keep  the  conditions  practically  constant.  The  personal 
factor  of  the  workmen,  in  mechanical  furnaces,  is  largely  though  not 
entirely  eliminated.  Maintaining  proper  conditions  in  a  roasting  furnace 
requires  skill,  but  as  the  extremely  hard  labor  of  hand  rabbling  the  ore  is 
eliminated  in  mechanical  furnaces,  there  is  not  the  same  temptation  to 
slight  the  work. 

The  output  per  shift  is  very  much  larger  in  a  mechanical  than  in  a 
hand  furnace,  so  that  it  makes  it  possible  to  pay  the  men  better  and  get  a 
superior  quality  of  labor.  In  mechanical  furnaces,  the  principal  factor 
under  the  control  of  the  roasterman,  is  the  temperature  of  the  furnace, 
and  even  this  may  be  made  largely  automatic  by  the  judicious  use  of 
pyrometers.  The  principal  function  of  the  roasterman,  in  large  mechan- 
ical roasters,  is  that  of  an  overseer,  and  to  regulate  the  temperatures  as 
indicated  by  the  pyrometers. 

Wonderful  strides  have  been  made  in  roasting  and  in  mechanical 
roasting  furnaces  in  recent  years.  Roasting  can  no  longer  be  considered 
either  difficult  or  expensive  if  fuel  is  available  at  a  reasonable  price. 
Roasting  costs  of  ten  or  twenty  years  ago,  in  hand  furnaces,  or  even  in 
mechanical  furnaces,  are  now  obsolete,  and  the  future  will  see  still  further 



The  mechanical  difficulties  of  roasting  furnaces,  prevailing  some 
years  ago,  which  are  peculiar  to  mechanism  working  under  high  tem- 
peratures and  in  the  presence  of  dust,  have  been  practically  overcome 
in  all  of  the  successful  mechanical  furnaces  now  in  use.  It  is  not  unusual 
for  a  large  100-ton  mechanical  roaster  to  run  from  three  to  six  months 
without  a  single  shut-down  except,  perhaps,  the  stopping  of  the  mechan- 
ism for  some  minutes  to  change  the  rabbles;  but  this  does  not  interfere 
with  the  daily  output  of  roasted  ore.  It  is  not  unusual  for  a  furnace  to 
run  from  six  months  to  a  year  without  cooling  for  repairs. 

Roasting,  in  mechanical  furnaces,  is  better  and  cheaper  than  in  any 
type  of  hand  furnace,  whether  the  amount  roasted  is  from  5  to  10  tons 
a  day,  or  from  100  to  200  tons,  for  a  single  furnace.  The  capacity  of 
200  tons  has  not  yet  been  realized,  nevertheless,  there  are  no  mechanical 
or  chemical  difficulties  to  its  realization.  Roasting  furnaces  of  that 
capacity,  in  large  works  and  on  low  grade  sulphur  ores,  will  soon  be  an 
established  fact,  and  will  considerably  reduce  present  costs  of  roasting. 
It  is  not  now  unusual,  on  Cripple  Creek  ores  containing  from  1  to  3  per 
cent,  sulphur,  to  roast  125  tons  a  day  in  furnaces  designed  to  roast,  nor- 
mally, 100  tons.  Since  it  usually  takes  one  man  on  a  shift  to  attend  to  a 
mechanical  roaster,  whether  the  capacity  is  25,  50,  or  100  tons  a  day,  the 
saving  in  the  larger  units  is  manifest,  as  well  as  the  saving  in  fuel  and  other 
items.  If  the  furnaces  are  fired  with  oil,  or  with  well  designed  centralized 
gas  producer  plant,  one  man  on  a  shift  can  attend  to  several  furnaces, 
irrespective  of  their  size,  with  an  extra  man  occasionally  to  assist  in 
attending  the  producers  and  changing  the  rabbles. 

In  small  installations,  the  first  cost  of  a  mechanical  furnace,  over  a 
long  reverberatory,  is  not  usually  a  serious  item.  The  difference  need  not 
exceed  $3000  to  $3500  'and  the  cost  of  roasting  can  usually  be  reduced 
from  50  to  75  per  cent.  A  good  mechanical  furnace  to  roast,  say,  10 
tons  of  pyritic  concentrates  a  day,  or  25  to  30  tons  of  low  sulphur  silicious 
ore,  can  be  erected  for  about  $6000.  It  will  take  three  of  the  ordinary 
long  hand  reverberatories  to  do  the  same  Avork.  Three  such  furnaces 
would  cost  more  than  the  mechanical  furnace,  and  would  require  one  man 
on  a  shift  for  each  furnace,  making  nine  men  in  all;  whereas  the  mechan- 
ical furnace  would  require  only  one  man  on  a  shift,  or  three  men  in  all. 
As  the  amount  of  ore  roasted  per  day  becomes  larger,  the  difference  in 
cost  of  roasting,  between  the  mechanical  and  hand  furnaces,  becomes 
more  pronounced. 

The  variety  of  roasting  furnaces,  evolved  and  suggested,  have  been 
numerous.  The  practice  has  all  been  toward  greater  simplicity.  Of 
the  hand  reverberatories,  the  multiple  hearth,  for  roasting  ore  for  sub- 
sequent treatment  by  chemical  processes,  has  become  obsolete.  The 
"Long  Reverberatory,"   the   "Fortschauflungsofen"    of  the   Germans, 


has  supplanted  all  other  types  of  hand  furnaces,  and  proved  itself  the 
survival  of  the  fittest. 

The  furnaces  to  consider  most  seriously  in  the  treatment  of  copper, 
gold,  and  silver  ores  by  the  solvent  processes,  are: 

Hand  reverberatories, 
Mechanical  reverberatories. 
Revolving  cylindrical  furnaces. 
Muffle  furnaces. 

All  of  these  types  are  in  actual  and  successful  use.  Shaft  furnaces, 
like  those  of  the  Stetefeldt  type,  have  long  since  become  obsolete.  There 
are  none  now  in  operation,  and  it  is  questionable  whether  a  roasting 
furnace,  based  on  that  principle,  can  ever  be  devised  which  will  success- 
fully compete  with  the  various  types  of  mechanical  reverberatories.  The 
chemical  conditions  of  roasting  in  a  shaft  furnace  are  all  that  could  be 
desired,  but  the  physical  and  mechanical  difficulties  are  well-nigh 

Hand  Reverberatories. — A  hand  reverberatory  roasting  furnace, 
fired  with  solid  fuel,  consists  essentially  of  a  hearth,  a  fire-box,  a  bridge 
separating  the  hearth  from  the  fire-box,  a  reverberatory  arch  over  the 
hearth  and  which  reverberates  the  heat  and  flame  toward  the  hearth,  a 
flue,  and  means,  such  as  exhauster  or  chimney,  of  acquiring  a  draft 
through  the  furnace.  If  gas  or  oil  are  used  as  fuel,  the  fire-box  and 
bridge  may  be  dispensed  with,  and  the  gas  or  oil  injected  through  the 
side  walls  or  through  the  arch. 

A  hand  reverberatory  is  one  in  which  the  ore  is  stirred  and  advanced 
by  hand  labor;  in  a  mechanical  reverberatory  the  ore  is  stirred  and 
advanced  by  mechanical  means. 

The  hand  reverberatories  may  be  subdivided  into  two  general  classes, 
based  essentially  on  the  method  of  operation.     These  are: 

1.  Tlie  short  reverberatory,  in  which  the  ore  is  all  charged,  roasted, 
and  withdrawn,  in  successive  complete  operations. 

2.  The  long  reverberatory,  in  which  the  ore  is  charged  at  one  end, 
and  then  advanced  by  stages,  while  at  each  stage  another  charge  is  intro- 
duced and  one  withdrawn.  Several  charges  are,  therefore,  in  the  furnace 
at  the  same  time  and  each  going  through  its  cycle  of  treatment,  inde- 
pendent of  the  others. 

Short  Reverberatory. — This  type  of  reverberatory  is  used  only  in 
works  where  small  quantities  of  ore  are  treated.  These  furnaces,  while 
cheap  to  construct,  are  expensive  to  operate.  They  labor  under  the 
disadvantage  that  the  conditions  of  the  furnace  itself  change  as  the  ore 
progresses  in  the  roasting  operation;  while  in  the  long  reverberatory  the 
conditions  of  the  furnace  remain  practically  the  same  all  the  time,  but  as 



the  ore  progresses  in  the  roasting  it  is  advanced  against  the  purer  air  and 
more  highly  oxidizing  atmosphere. 

Figs.  1  and  2  show  section  and  plan  respectively  of  a  short  reverbera- 
tory,  which  is  the  usual  form  and  construction  for  furnaces  of  that  size. 
It  has  a  hearth  area  of  approximately  120  sq.  ft.,  and  is  capable  of  roast- 
ng  a  ton  of  ore  at  a  charge.  The  number  of  charges  that  can  be  roasted 
in_24  hours  depends  upon  the  ore;  if  the  ore  is  silicious  and  low  in  sulphur, 
three  charges  a  day  can  be  roasted;  if  the  ore  consists  of  pyritic  concen- 
trates, from  one  to  two  charges  a  day  is  about  all  that  could  be  put 
through.  If  the  hearth  is  made  longer  than  12  ft.,  it  is  better  to  throw 
the  arch  longitudinally  over  the  hearth,  instead  of  transversely,  as  shown. 

In  Figs.  1  and  2,  A  represents  the  hearth,  B  the  reverberatory  arch, 
C  the  fire-box,  D  the  bridge  wall,  E  the  flue  holes  leading  from  the  rever- 
beratory chamber  to  a  small  fine  chamber  before  entering  the  stack, 


Fig.   1. — Short  hand  reverberatory.     Longitudinal  section. 

F  the  charging  hole  through  the  arch  from  the  top  of  the  furnace,  and 
H  a  hole  through  the  hearth  for  discharging  the  roasted  ore  through  a 
spout  on  to  the  cooling  floor,  or  into  a  car  or  wheel-barrow,  to  be  taken 
to  the  cooling  floor.  The  flues,  E,  are  arranged  so  that  the  flame  from 
the  fire  can  be  equally  distributed  over  the  entire  body  of  ore  on  the 
hearth.  The  holes,  K,  in  the  outer  wall,  easily  admit  of  the  regulation  of 
the  flue  holes  by  means  of  bricks  placed  in  the  flues.  The  stack,  also, 
should  be  provided  with  a  suitable  damper. 

It  is  not  necessary  to  build  the  entire  interior  of  the  furnace  of  fire 
brick.  The  fire-box,  bridge,  and  arch  immediately  over  them  must  be 
built  of  fire  brick;  the  rest  may  be  built  of  any  good  common  red  brick, 
preferably  a  pressed  brick  of  the  cheaper  quality.  Common  pressed 
brick  for  the  hearth  is  very  desirable;  it  is  hard  enough  to  withstand  the 
wear,  and  smooth  and  even  enough  to  make  the  rabbling  easier  than  it 



would  be  otherwise.  The  hearth  brick  should  be  set  on  edge,  4  in.  thick, 
and  laid  without  mortar.  The  joints  may  afterward  be  filled  in  with 
fine  sand  or  tailings. 

On  account  of  the  bridge  being  exposed  to  injury  by  the  high  tempera- 
ture on  one  side,  and  by  rabbling  on  the  other,  it  is  desirable  that  the  top 
course  should  be  made  of  fire  clay  tile,  say,  12  in.  by  24  in.  by  2  in.  thick. 
The  grates  are  12  to  16  in.  below  the  top  of  the  bridge,  and  the  top  of  the 
bridge  is  8  to  10  in.  above  the  hearth.  Through  the  middle  of  the  arch 
is  an  opening  of  cast  iron,  F,  with  a  well-fitted  cast-iron  cover,  through 
which  the  ore  is  charged  into  the  furnace.  The  walls  of  the  furnace 
should  be  reasonably  thick,  in  order  to  retain  the  heat  as  much  as  possible. 

Fig.  2. — Short  hand  reverberatory.    Plan. 

and  to  prevent  the  furnace-room  from  getting  uncomfortably  hot.  The 
arch  should  be  at  least  8  in.  thick,  with  brick  set  on  edge.  Another  4-in. 
arch  may  be  placed  on  top  of  this,  if  desired.  Most  of  the  heat  radi- 
ated from  a  furnace  passes  through  the  arch,  so  that  the  cost  of  the  extra 
thickness  of  arch  is  money  well  invested.  For  the  arch,  fire  clay  should  be 
used  for  the  joints,  and  the  joints  are  best  made  by  dipping  the  brick 
into  the  thin  clay  before  setting  them  in  place.  If  a  4-in.  arch  is  placed 
over  the  8-in.  arch,  it  may  be  made  of  brick  or  brickbats,  laid  in  mortar. 
The  rabble  doors  are  usually  about  8  in.  high  and  14  to  16  in.^  wide.  In 
front  of  the  door  is  an  iron  bar  laid  across  it  from  projections  on  the 
■casting,  to  facilitate  the  rabbling.  The  furnace  is  bound  together  by 
1-in.  rods,  attached  either  to  cast  iron,  or  railroad  iron  buck  staves. 

Long  Reverberatory. — The  typical  hand  roasting  furnace  is  the  long 
reverberatory,  of  which  Figs.  3,  4,  5,  6,  and  7  show  a  typical  example. 
It  is  the  type  of  hand  reverberatory  almost  universally  used,  and  is 


r  '  I  n  //    ' 














essentially  the  same  as  those  used  in  smelting  works.  The  length  of  the 
furnace  is  largely  governed  by  the  character  of  the  ore  to  be  roasted.  If 
the  ore  is  highly  silicious,  and  contains  only  a  small  amount  of  sulphur, 
there  is  no  advantage  in  having  the  length  more  than  40  to  50  ft.  If  the 
ore  is  highly  pyritic,  or  consists  of  ordinary  sulphide  concentrates,  a 
length  of  60  to  75  ft.  will  be  found  the  most  satisfactory.  With  furnaces 
longer  than  60  to  75  ft.,  it  will  be  found  difficult  to  ignite  the  ore  at  the 
rear,  and  unless  it  is  ignited,  except  for  drying  and  heating,  there  is  no 
advantage  in  having  the  furnace  much  longer  than  the  zone  of  ignition. 
This  zone,  under  any  conditions,  is  considerably  shorter  in  furnaces 
roasting  for  the  chemical  processes  than  for  smelting,  because  the  initial 
heat  at  the  fire-box,  and  hearth  adjoining,  must  always  be  kept  below 
the  sintering  point  of  the  ore,  whereas  if  the  ore  is  roasted  for  smelting, 


mtijjj  uujj  I  u  uj  u  ujjj  u  ti  ijjjjmujjjiiiJJiJiJJiD ' jjjjj  ouJiUUJij  i  ijiii':  u 

— 16-> 

I'lG.  5. — Long  hand  reverberatory.     Transverse  section. 

the  sintering  is  of  no  particular  consequence.  It  will  usually  be  found 
advisable,  therefore,  if  increased  capacity  is  desired,  to  build  two  furnaces, 
rather  than  to  attempt  to  get  it  by  increasing  the  length  of  the  hearth, 
or  perhaps  by  the  addition  of  a  second  fire-box,  toward  the  rear. 

The  practical  width  of  the  furnace  is  controlled  by  the  convenience  of 
working  from  both  sides.  From  14  to  16  ft.  has  been  found  by  experi- 
ence to  give  the  best  general  results.  When  the  width  exceeds  16  ft.,  or  a 
reach  of  rabbling  of  about  8  ft.,  the  labor  of  stirring  the  ore  and  advancing 
it  becomes  tiresome  for  the  workmen,  and  hence  the  quality  of  the  roast 
is  likely  to  be  defective. 

The  hearth  is  sometimes  made  continuous,  in  one  plane,  and  some- 
times with  slight  breaks  of  several  inches,  corresponding  in  length  of 
hearth  to  the  amount  of  ore  charged  into  the  furnace  at  one  time.  The 
object  of  the  stepped  hearth  is  to  enable  the  roastermen  to  clearly  dis- 
tinguish the  different  charges,  and  keep  them  separate  as  they  progress 
through  the  furnace.  As  one  charge  is  sufficiently  roasted  and  withdrawn, 
the  next  charge  is  moved  forward  to  the  position  occupied  by  the  previous 



one,  and  a  new  charge  introduced  in  the  section  at  the  flue  end.  It  is 
very  desirable  that  mixing  of  the  different  charges  should  not  occur.  If 
mixing  occurs  to  any  considerable  extent,  the  insufficiently  roasted  ore 
will  contaminate  that  which  is  well  roasted  and  serious  difficulty  in  the 
chemical  treatment  will  be  the  result.  If  the  hearth  is  not  stepped,  the 
roastermen  work  to  imaginary  lines  in  the  furnace  to  keep  the  charges 
separate.  Sometimes,  in  order  to  enable  the  workmen  to  more  easily 
advance  the  ore,  the  hearth  is  built  with  a  gentle  slope  from  rear  to 
front.  The  hearth  should  be  built  4  in.  thick  with  brick  laid  on  edge 
and  without  mortar.  The  ordinary  quality  of  pressed  brick  make  an 
excellent  hearth,  and  these  are  satisfactory  for  the  arch  also. 

Fig.  6. 

Fig.  7. 
Figs.  6  and  7. — Long  hand  reverberatoiy.     Details  of  construction. 

The  height  of  the  arch  above  the  hearth  is  dependent  largely  upon 
the  nature  of  the  ore  to  be  roasted.  As  a  matter  of  fuel  economy,  the 
lower  the  arch,  the  more  flame  and  fuel  gases  will  come  directly  in  con- 
tact with  the  ore,  but  the  limit  in  this  direction  is  governed  by  the  condi- 
tions of  the  furnace  atmosphere.  When  roasting  highly  silicious  ores, 
the  furnace  gases  will  be  highly  oxidizing  under  almost  any  conditions, 
and  the  height  of  the  arch  above  the  hearth,  in  such  a  case,  is  therefore 
limited  by  other  considerations,  such  as  convenience  in  rabbling.  If, 
however,  the  ore  is  high  in  sulphur,  as  in  roasting  pyritic  concentrates, 
the  fumes  of  sulphur  dioxide  from  the  ore  and  carbon  dioxide  from  the 


fuel,  would  tend  to  make  the  furnace  atmosphere  reducing,  instead  of 
highly  oxidizing,  and  the  object  of  roasting  would  be  largely  defeated. 
With  such  ores,  the  reverberatory  arch  should  be  high  so  as  to  permit  of 
large  volumes  of  air  passing  over  the  ore  in  order  to  keep  the  atmosphere 
within  the  furnace  as  highly  oxidizing  as  possible.  From  20  to  36  in. 
will  usually  be  found  to  be  within  the  practical  limits  of  the  height  of  the 
crown  of  the  arch  above  the  hearth.  The  rise  of  the  arch  should  not  be 
less  than  3/4  to  1  in.  for  every  horizontal  foot  of  width.  That  is  to  say, 
the  least  rise  that  an  arch  16  ft.  wide,  should  have  to  be  safe,  is  from  12 
to  16  in.  Theoretically,  a  flat  arch  is  the  best,  but  there  are  practical 
difficulties  in  constructing  and  maintaining  a  flat  arch  under  the  strenuous 
conditions  to  which  it  would  be  subjected  in  furnace  work.  So  far  as  the 
roasting  is  concerned,  the  best  results  are  obtained  by  having  the  arch  as 
nearly  flat  as  possible,  and  the  limit  in  this  direction  is  governed  by  con- 
structional difficulties.  As  to  the  best  practical  rise  for  the  reverberatory 
arch,  much  will  depend  on  the  quality  of  the  brick;  but  the  rule  given 
above  will  be  found  best  under  average  conditions  and  conform  with  the 
best  practice  for  all  types  of  reverberatories  for  roasting  ores. 

The  method  of  constructing  the  arch  is  largely  a  matter  of  choice. 
The  brick  should  be  set  on  edge,  the  8-in.  way.  The  arch  may  be  built 
of  independent  successive  rings,  as  shown  in  Fig.  8,  or  be  bonded  so  as  to 
make  a  continuous  whole,  as  shown  in  Fig.  9.  The  method  of  independ- 
ent rings  has  much  to  commend  it.     With  this  construction,  every  brick 

Fig.  8.  Fig.  9. 

Figs.  8  and  9. — Reverberatory  arch  construction. 

is  under  full  compression,  whereas  in  the  bonded  arch,  if  the  brick  or 
joints  are  not  of  equal  size,  it  is  not  possible  to  have  them  all  under  the 
same  compression,  so  that  there  is  danger  of  the  thinner  ones  dropping 

The  reverberatory  arch  is  usually  built  of  ordinary  straight  brick, 
with  the  difference  in  the  thickness  of  the  joint  between  the  intrados  and 
extrados  made  up  with  clay.  It  is  better,  however,  to  build  the  arch  of 
straight  brick  and  occasionally  insert  a  row  of  wedge  brick.  In  this  way 
the  joints  can  all  be  kept  of  even  thickness  and  the  arch  will  have  greater 
stability.  It  will  not  be  necessary  to  carry  the  fire  brick  in  the  arch 
more  than  10  or  20  ft.  beyond  the  fire-box,  in  the  roasting  chamber. 


The  remainder  of  the  arch,  as  also  the  side  walls,  may  be  built  of  common 

The  side  and  end  walls  should  not  be  less  than  16  in.  thick,  or  two 
courses  of  brick  laid  the  8-in.  way.  A  little  extra  cost  in  thick  walls  and 
reverberatory  arch,  will  be  more  than  compensated  for  in  the  saving  of 
fuel  and  comfort  to  the  workmen. 

The  foundations  are  best  and  cheapest  made  of  concrete.  By  putting 
up  the  necessary  side  boards,  and  leveling  the  top  edges,  the  foundation 
can  be  quickly  and  cheaply  laid  in  the  best  possible  condition  for  the 
superstructure.  The  concrete  may  be  brought  up  to  within  a  foot  of  the 
hearth  level,  and  this  also  will  be  found  cheaper  than  brickwork  and  be 
just  as  good.  Brickwork,  in  mining  districts,  and  especially  in  isolated 
camps,  is  expensive,  since  all  the  material  has  to  be  hauled  on  the  ground, 
and  brick-layers  command  high  wages.  With  concrete,  only  the  cement, 
which  is  a  small  proportion  of  the  whole,  has  to  be  supplied  from  without, 
since  rock  and  sand  are  usually  available  and  common  labor  all  that  is 

The  lower  tie-rods  should  not  be  built  in  solid,  but  ducts  should  be 
provided  so  that  they  can  be  removed  or  inserted  at  will,  should  the  rods 
at  any  time  become  disabled.  Worn  out  iron  piping,  or  common  earthen- 
ware pipes,  are  the  best  for  this  purpose,  although  ducts  made  of  brick 
will  answer  about  as  well. 

The  space  between  the  walls,  below  the  hearth,  may  be  filled  with  rock 
and  earth  or  loam.  It  should  be  well  tamped,  so  that  there  will  be  no 
danger  of  the  hearth  settling  when  the  furnace  is  in  operation. 

The  rabble  door  frames  should  be  set  flush  with  the  hearth  and  also 
with  the  exterior  walls.  The  buckstaves  will  then  lap  the  j  oint  between 
the  exterior  brickwork  and  the  iron  castings,  and  catch  the  face  of  the 
channel  which  takes  the  thrust  of  the  arch.  The  rabble  door  castings 
are  in  this  way  securely  fastened  without  bolts.  Instead  of  the  channel 
as  shown  in  the  drawings,  cast  iron  beams  in  the  form  of  T's  may  be  used 
to  support  the  arch,  and  which  alternate  with  the  rabble  door  castings. 
The  buckstaves  may  be  made  of  old  railroad  rails,  I-beams,  cast  T-irons 
or  two  small  channels  secured  together,  back  to  back,  with  a  separator  to 
permit  the  tie-rods  to  go  between  them.  It  is  well  to  have  the  tie-rods 
abundantly  heavy  to  take  the  enormous  horizontal  thrust  of  the  arch. 
For  the  ordinary  span  of  from  12  to  16  ft.,  1  1/4-in.  rods  have  been  found 
satisfactory.  For  the  longitudinal  rods  3/4-in.  diameter  will  be  large 

The  number  of  working  doors  should  be  sufficiently  numerous  to 
permit  of  easy  rabbling.  A  distance  of  about  6  ft.,  from  center  to  center, 
is  satisfactory.  When  the  doors  are  too  far  apart,  rabbling  of  the  ore  in 
the  intermediate  spaces  becomes  difficult,  and  may  be  neglected.  Figs. 
6  and  7  show  the  details  of  the  working  doors  and  the  method  of  setting 


them.  The  details  may  vary  somewhat  according  to  the  material  most 
convenient,  and  upon  local  conditions.  Each  door  casting  has  attached 
to  it  lugs  which  receive  an  iron  bar  about  an  inch  square  upon  which  the 
rabble  may  slide  while  the  ore  is  being  rabbled.  These  bars  are  remov- 
able, and  sometimes  it  is  more  convenient  to  work  without  them.  The 
door  castings  have  flanges  on  the  sides  so  that  they  can  be  secured  in 
place  by  the  buckstaves.     Each  door  casting  weighs  about  150  lb. 

When  odd  shaped  bricks  are  used,  it  will  be  found  cheaper  and 
better  to  have  them  especially  made  for  the  purpose  if  they  cannot 
be  obtained  in  standard  shapes.  The  cutting,  and  consequent  breaking, 
of  a  large  number  of  brick  will  be  more  expensive  than  the  extra  cost  of 
special  forms,  and  the  work  will  not  be  as  good.  Usually  the  manufac- 
turers of  fire  brick  have  special  shapes  enough  to  fill  any  want. 

All  brickwork  about  a  furnace  should  be  "shoved"  and  well  grouted, 
especially  that  part  of  the  work  which  forms  the  skewback  of  the  rever- 
beratory  arch.  The  spandrels  of  the  arch  may  be  filled  in  with  brickbats 
and  mortar.  The  filling  in  of  the  spandrels  will  make  the  arch  stronger 
and  also  enable  the  furnace  to  better  retain  the  heat. 

After  the  furnace  is  finished,  the  buckstaves  and  tie-rods  should  be 
put  in  place.  The  rods  should  be  made  reasonably  tight,  so  that  they 
will  vibrate  when  struck.  It  is  best  to  take  up  any  looseness  in  the  arch 
l)efore  its  weight  is  taken  off  the  centers.  The  centers  may  then  be  lo- 
moved,  either  before  firing,  or  they  may  be  burned  out  afterward,  which- 
ever is  the  most  convenient  and  economical.  The  furnace  should  be 
allowed  to  stand  as  long  as  possible  before  firing.  When  the  fire  is 
started  the  furnace  should  be  heated  quite  gradually  for  at  least  24  hours, 
after  which  there  will  be  no  harm  in  bringing  it  up  to  heat.  It  should  be 
fired  long  enough  to  get  the  hearth  quite  hot  before  charging  the  ore, 
since  a  cold  hearth  greatly  retards  the  roasting. 

The  furnace  should  be  provided  with  several  rabbles,  4  in.  by  8  in., 
and  12  to  14  ft.  long,  and  several  paddles  8  in.  by  12  in.,  14  ft.  long.    The 
handles  are  best  made  of  strong  wrought  iron  pipe  to  which  the  rabble 
'  and  paddle  blades  are  fastened,  by  welding. 

The  stack  of  the  furnace  may  be  built  of  brick,  iron,  or  re-enforced  con- 
crete. For  single  furnaces  an  iron  stack  will  usually  be  found  the  most 
economical.  The  stack  for  a  furnace  as  shown  in  Figs.  3,  4,  and  5,  should 
be  about  30  in.  in  diameter  and  from  60  to  75  ft.  high.  The  stack,  or 
the  flue  leading  to  the  stack,  should  be  provided  with  a  damper.  The 
position  of  the  stack  in  reference  to  the  furnace,  is  largely  a  matter  of 
convenience  and  local  conditions.  If  the  works  have  several  furnaces, 
they  may  all  connect  with  a  common  stack.  If  there  is  only  one  furnace, 
and  no  dust  chamber  is  desired,  the  arrangement  of  the  stack  as  shown 
in  Figs.  3  and  4  will  be  as  satisfactory  as  any.  Provision  should  be  made 
for  a  car  track  to  bring  the  ore  to  the  hopper  over  the  rear  of  the  furnace. 


Hand  reverberatories  should  not  make  more  than  3/4  to  1  1/2  per 
cent.  dust.  Whether  this  dust  is  worth  recovering  by  building  a  large 
dust  chamber  will  depend  largely  on  the  value  of  the  ore.  Small  dust 
chambers  are  not  very  effective.  It  will  ordinarily  be  found  that  the 
recovery  of  the  dust  is  a  matter  worth  careful  consideration. 

The  cost  of  building  a  long  reverberatory,  as  described  and  shown, 
will  be  between  $3000  and  14000.  It  will  take  about  50,000  common 
brick,  8000  fire  brick,  and  20,000  lb.  of  iron.  If  concrete  is  largely  used, 
a  great  saving  may  be  effected  in  the  number  of  brick,  and  a  saving  also 
in  the  cost  of  the  furnace. 

Method  of  Operating  a  Long  Hand  Reverberatory. — When  the  furnace 
is  hot,  the  ore,  which  may  be  assumed  to  be  pyritic  concentrates,  is 
charged  from  a  car  into  the  hopper,  and  into  the  furnace,  in  the  section 
nearest  the  flue,  or  at  the  rear.  It  is  then  spread  out  evenly  over  this 
section  of  the  hearth.  The  weight  of  the  charge,  for  the  best  work,  should 
not  exceed  10  to  15  lb.  per  square  foot  of  hearth  area;  10  lb.  is  better  than 
15  lb.  if  the  ore  contains  very  much  sulphur.  In  any  event,  there  is  no 
advantage  in  using  a  deep  bed  of  ore,  for  what  is  gained  in  the  amount  of 
ore  charged,  is  lost  by  a  correspondingly  increased  time  of  roasting.  The 
depth  of  the  charge  will  usually  be  from  2  to  3  in. 

The  working  doors  are  all  closed  until  the  sulphur  is  well  ignited.  The 
moisture  is  first  driven  off,  after  which  the  ore  will  soon  become  slightly 
incandescent  and  the  sulphur  begin  to  burn  with  a  blue  flame.  This  is 
one  of  the  most  delicate  stages  of  the  roasting,  and  should  be  done  at  the 
lowest  permissible  temperature  and  in  the  presence  of  a  maximum 
.  amount  of  air.  The  charge  should  be  rabbled  energetically,  and  with  the 
intervals  between  the  rabbling  as  short  as  possible.  At  this  stage  the  ore 
will  be  very  unstable  and  is  likely  to  run  somewhat  like  a  liquid.  If  the 
rabbling  is  neglected,  or  if  the  temperature  of  the  furnace  is  too  high, 
partial  fusing  or  matting  is  likely  to  occur,  which  forms  lumps  that  are 
difficult  to  eliminate;  and  if  not  eliminated,  will  result  in  improperly  and 
insufficiently  roasted  ore.  Such  ore  will  be  highly  detrimental  in  the 
subsequent  chemical  treatment.  The  excess  of  air  required  during  this 
stage  of  the  roasting  may  be  obtained  by  keeping  the  working  doors  open, 
and  by  free  admission  of  air  through  the  bridge  wall.  The  air  should  be 
introduced  as  much  as  possible  through  the  bridge,  since  this  will  tend  to 
keep  the  charge  nearest  the  fire  from  becoming  too  hot,  by  interposing  a 
layer  of  cooler  air  between  the  ore  and  the  flame,  and  give  a  highly  oxidiz- 
ing atmosphere  at  the  surface  of  the  ore  where  it  is  most  desired. 

When  the  sulphur  flame  has  abated,  which  will  be  in  about  8  hours 
after  charging,  the  ore  is  moved  forward  into  section  No.  2,  and  a  new 
charge  introduced  into  section  No.  1.  The  ore  on  section  No.  2  is  spread 
out  over  a  large  area  to  give  it  as  much  surface  as  possible.  There  is 
still  much  sulphur  in  the  ore,  and  most  of  the  oxidation  takes  place  in  the 


middle  section.  The  ore  being  brought  closer  to  the  fire,  is  brought  to  a 
dull  red  heat.  During  this  stage  the  ore  swells  somewhat,  and  becomes 
more  or  less  inert.  As  the  sulphur  is  eliminated,  the  ore  has  no  power  of 
generating  heat  within  itself  and  hence  the  fire  is  urged,  to  keep  the  ore 
at  the  desired  temperature.  The  rabbling  in  this  section  need  not  be  as 
frequent  as  in  section  No.  1;  a  thorough  stirring  every  15  or  20  minutes 
will  suffice.  The  ore  should  be  uniform  throughout,  and  as  it  is  turned 
over,  the  newly  exposed  incandescent  surface  should  quickly  turn  dark 
and  not  show  any  live  sparks  of  burning  sulphur. 

The  charge,  after  reaching  this  stage,  which  will  usually  be  about  16 
hours  after  it  has  been  put  into  the  furnace,  is  transferred  to  section  No.  3; 
the  ore  in  No.  1  is  advanced  to  section  No.  2,  and  a  new  charge  introduced 
into  section  No.  1.  There  is  now  no  danger  of  lumps  forming  in  section 
No.  3,  and  the  temperature  may  be  raised  somewhat,  but  must  never 
approach  the  sintering  point.  The  temperature  permissible  in  this 
section  is  the  controlling  factor  in  firing  the  furnace.  The  firing  should 
always  be  done  with  a  view  of  throwing  as  much  heat  and  flame  as  possible 
to  the  rear  of  the  furnace  without  danger  of  sintering  the  ore  on  the 
finishing  section.  In  this  section,  the  roasting  will  largely  consist  of 
decomposing  the  soluble  sulphates,  and  while  this  is  going  on  the  odor  of 
sulphur  dioxide  can  be  detected  when  a  sample  of  the  incandescent  ore 
is  removed  for  inspection.  As  the  ore  becomes  more  nearly  completely 
roasted,  it  becomes  more  coherent,  and  remains  as  placed  by  the  rabble. 

"Sweet"  or  "Dead"  roasting  are  more  or  less  indefinite  terms,  used 
to  denote  the  condition  of  the  ore  when  all  the  sulphur  has  been  elimi- 
nated. But  as  the  elimination  of  all  the  sulphur  is  practically  impossible, 
and  as  the  approximation  thereto  is  a  very  indefinite  matter  varying  with 
the  different  ores,  it  may  be  taken  to  mean  ore  sufficiently  roasted  to  give 
the  best  results  in  the  subsequent  chemical  treatment. 

After  the  ore  has  remained  on  the  finishing  section  for  about  7  hours, 
the  roasting  is  completed.  It  is  then  withdrawn  through  the  holes  in  the 
hearth,  near  the  last  door,  into  a  pit,  or  into  a  car  and  taken  to  the  cooling 
floor.  The  charge  in  section  No.  2  is  then  moved  forward  to  No.  3,  and 
the  charge  in  section  No.  1  moved  forward  to  No.  2,  while  a  new  charge  is 
introduced  into  section  No.  1.  In  this  way  a  charge  is  withdrawn  and  a 
new  one  added  every  day,  so  that  there  are  always  three  separate 
charges  in  the  furnace,  and  each  charge  remains  in  the  furnace  almost 
24  hours. 

The  fuel  used  in  roasting  should  be  either  wood  or  long  flame  coal. 
Oil  gives  better  results  than  either  wood  or  coal,  but  is  not  usually  avail- 
able. It  is  best,  in  order  to  get  a  long  flame,  with  almost  any  coal,  to 
fire  the  furnace  fire-box  more  or  less  as  a  gas  producer.  This  is  easily 
arranged  by  keeping  a  deep  bed  of  ash  and  fuel  on  the  grates,  and  intro- 
ducing steam  and  air  through  the  closed  ash  pit.     Much  of  the  air  needed 


to  completely  consume  the  fuel  gases  may  be  introduced  through  the 
bridge  and  some  through  the  working  doors.  The  draft  is  regulated  by 
the  damper  in  the  flue  or  stack,  and  by  opening  and  closing  the  working 
doors.  The  time  of  roasting  depends  largely  on  the  ore,  but  somewhat 
also  on  the  amount  of  rabbling.  Roasting,  as  already  stated,. is  essen- 
tially an  oxidizing  process,  so  that  any  operation,  such  as  continuous  and 
energetic  rabbling,  which  will  expose  the  greatest  amount  of  ore  to  the 
highly  oxidizing  furnace  atmosphere,  will  materially  reduce  the  time  of 

It  is  customary  when  roasting  pyritic  gold  concentrates,  or  ore  con- 
taining lime,  or  silver  and  copper  in  appreciable  quantities,  to  add* a  small 
amount  of  salt,  usually  just  before  drawing  the  charge.  The  amount  of 
salt  may  vary  from  0.5  to  5  per  cent.  If  the  ore  is  not  sensitive  to  vola- 
tilization, the  salt  may  be  added  to  the  ore  as  it  is  advanced  from  the 
middle  to  the  finishing  section  and  thus  become  thoroughly  incorporated 
with  it.  Usually,  however,  the  salt  is  added  about  30  minutes  before  the 
charge  is  withdrawn,  and  thoroughly  mixed  with  the  ore.  Chloridization, 
under  proper  conditions,  takes  place  rather  quickly,  and  as  explained 
under  " Chloridizing  Roasting"  air  is  not  essential  to  the  chloridization. 
The  salt  may  therefore  be  added  a  short  time  before  discharging  the  ore, 
and  by  permitting  the  ore  to  cool  slowly  after  it  is  discharged,  the  neces- 
sary degree  of  chloridization  can  well  be  realized  without  any  appreciable 
loss  by  volatilization.  This  gives  the  ore  a  thorough  oxidizing  roast 
before  chloridization,  and  the  sulphur  is  never  so  thoroughly  eliminated 
but  that  there  are  always  enough  sulphates  left  to  sufficiently  chloridize 
the  silver  and  small  amounts  of  copper. 

After  the  salt  is  added,  the  ore  begins  to  fume,  increase  in  bulk,  and 
has  a  "woolly"  appearance.  After  the  salt  is  added  the  temperature 
should  be  kept  low — not  over  a  dull  red  heat.  Much  of  the  gold  is 
chloridized  as  well  as  the  silver;  careful  tests  have  shown  it  to  be  from  10 
to  20  per  cent,  of  the  gold  contained  in  the  ore. 

If  the  ore  contains  galena,  great  care  must  be  exercised  in  the  first 
stages  of  the  roasting  to  keep  the  charge  at  the  lowest  practicable  tem- 
perature, as  the  lead  sulphide  fuses  at  a  very  low  heat,  and  agglomeration 
in  the  early  stages  of  the  roasting  will  make  the  subsequent  work  more 

Furnaces  having  considerable  of  a  drop  between  the  different  sections 
of  the  hearth  have  been  recomended  and  built,  but  have  not  come  into 
general  use.  The  cause  of  this  is  evident.  While  theoretically  the 
showering  of  ore  through  this  drop,  as  in  a  shaft  furnace,  appears  good,  it 
is  evident  that  the  draft  in  the  furnace  will  whip  the  dust  along  with  it 
and  cause  excessive  loss  in  that  way.  The  great  dust  loss  is  not  com- 
pensated for  by  the  small  gain  in  the  time  of  roasting. 

Most  furnaces,  like  the  one  illustrated  in  Figs.  3,  4,  and  5,  are  de- 


signed  to  take  a  charge  of  from  3  to  3  1/2  tons  on  a  section.  It  is  usual, 
however,  instead  of  charging  and  withdrawing  this  amount  of  ore  all  at 
once,  to  still  furthur  subdivide  it  so  that  each  shift  of  eight  hours  will 
charge  and  withdraw  one  third  of  this  amount  or  from  2000  to  2400  lb. 
The  advantage  of  this  is  that  the  quality  and  quantity  of  the  roasted 
material  can  be  checked  up  for  the  different  shifts,  and  the  furnace  can 
be  worked  with  greater  regularity,  than  when  so  much  ore  is  charged  and 
withdrawn  at  the  same  time. 

One  man  on  a  shift,  working  three  shifts,  will  roast  from  3  to  3  1/2 
tons  of  pyritic  concentrates  a  day.  Each  shift  will  draw  a  charge  of 
from  2000  to  2400  lb.  and  introduce  one.  The  amount  of  fuel  used  is  1/2 
cord  of  wood  per  ton  of  concentrates. 

The  following  tabulated  statement  gives  the  essential  facts  of  roast- 
ing pyritic  concentrates  in  California:' 


Eureka  and  Idaho 
mines,  Graas  Valley 










Washington  mine, 
Mariposa  County 










Black  Bear  mine, 
Klamath  County 














Oxygen  and  loss  by  difference .  . 


These  analyses  give  a  very  good  idea  of  the  composition  of  California 
pyritic  concentrates,  which  have  been  treated  for  many  years  by  roasting 
and  chlorination.  The  size  of  some  of  the  furnaces,  the  time  of  working 
on  a  shift,  and  the  quantity  of  ore  treated,  is  given  in  the  following  table: 

Name  of  works 

Size  of  furnace 

Time  of  working  (shift)         Quantity  of  ore  in  24  hours 



Amador. . . 
Merri  field's 


'  Eggleston, 

'Met.  Silver,  Gold,  Mer. 

8  hours 
8  hours 
8  hours 
8  hours 
8  hours 
12  hours 

3  tons 

3  tons 

4  1/2  tons 
3  tons 

2  1/2  tons 

3  tons 








The  hearth  of  these  furnaces  lasted  from  four  to  six  years. 

Cost  of  Roasting  in  Long  Reverberatories.— The  cost  of  roasting  in 
long  reverberatories  is  quite  large.  For  pyritic  concentrates,  which  is 
about  the  only  material  roasted  in  these  furnaces,  it  is  about  $4.25  per 
ton,  distributed  as  follows: 

Roasterman,  1  shift,  roasting  one  ton,  $2 .  50 

Fuel,  1/2  cord  of  wood,  for  one  shift,  1 .  50 

Other  expenses,  0  25 


The  cost  of  roasting  silicious  ore,  low  in  sulphur,  is  very  much  less. 
Such  ore  can  be  roasted  about  as  rapidly  as  it  can  be  worked  through  the 

In  California'  at  one  of  the  mills  where  oil  was  substituted  for  wood, 
it  was  found  that  the  capacity  of  the  furnace  was  increased  from  4  tons 
to  6  tons  per  day.  The  furnace  was  14  ft.  wide  and  75  ft.  long.  Bakers- 
field  crude  oil,  of  14  to  16°  gravity,  was  used.  In  roasting  2647  1/4  tons 
of  pyritic  concentrates,  1290  barrels  of  Bakersfield  crude  oil  was  used 
which  cost,  delivered,  $1917.63  or  48/100  barrel  per  ton  of  ore,  cost- 
ing 72  cents.  There  was  also  used  66.76  tons  of  coal  to  generate  steam 
for  pumping,  heating,  and  atomizing  the  oil,  which  cost  delivered, 
1867.84,  or  35  cents  per  ton  of  ore  treated;  making  a  total  cost  of  fuel  for 
roasting,  of  $2785.47  or  $1.05  per  ton. 

Fig.  12. — Modified  long  hand  reverberatory.     Transverse  section. 

Modified  Long  Reverberatories. — An  important  modification  of  the 
long  reverberatory,  especially  for  copper  roasting,  is  shown  in  Figs.  10, 11, 
and  12.  Fig.  10  shows  a  longitudinal  section;  Fig.  11  the  plan,  and 
Fig.  12  a  transverse  section  through  the  front  of  the  furnace  at  the  pro- 
tecting arch  and  discharge  openings  in  the  hearth. 

One  of  the  difiiculties  in  hand  reverberatories  in  roasting  ore  for  the 
hydrometallurgical  processes  is  the  liability  to  fuse  the  ore  near  the  fire 

'  E.  C.  Vorhies,  Scientific  and  Mining  Press,  March  26,  1904. 



end,  when  the  fire  is  urged  sufficiently  to  ignite  the  charge  at  the  rear. 
In  chloridizing  roasting  a  similar  difficulty  presents  itself.  In  order  to 
eliminate  the  sulphur  sufficiently  in  hearth  section  No.  2,  by  oxidizing 
roasting,  a  temperature  nearly  as  high  as  the  ore  will  stand  without  fusing 
is  desirable  to  expedite  the  process  as  much  as  possible.  To  accomplish 
this  the  ore  in  section  No.  3  may  become  unduly  heated,  even  if  it  does 

Fig.  13. — Modified  hand  reverberatory.     Longitudinal  section. 

not  approach  the  sintering  point.  The  temperature  for  chloridizing 
roasting  should  be  the  lowest  at  which  the  reactions  take  place,  and  this 
condition  conflicts  with  that  required  for  efficient  roasting  in  the  middle 
and  rear  end  of  the  furnace. 

To   overcome   these   difficulties,    a  protecting  arch  is  thrown  over 
section  No.  3,  as  shown  in  Figs.  10  and  12,  which  shields  the  ore  from  the 

Fig.  14. — Modified  hand  reverberatory.     Plan. 

direct  action  of  the  heat  and  flame.  In  this  way  the  ore  in  the  middle 
section  may  be  kept  the  hottest,  and  that  in  the- rear  section  may  readily 
be  ignited.  The  flame  and  heat  from  the  fire-box  pass  through  the 
space  between  the  protecting  arch  and  the  reverberatory  arch,  so  that 
the  ore  in  the  finishing  section  is  heated  only  indirectly,  as  in  a  muffle. 
The  protecting  arch  should  be  as  thin  as  possible,  consistent  with  good 



In  the  chloridizing  roasting  of  copper  ores,  this  modified  furnace  has 
been  used  in  preference  to  the  ordinary  long  reverberatory.  It  can 
readily  be  seen,  that  for  chloridizing  work  especially,  this  modification  of 
the  long  reverberatory  offers  many  advantages.  The  furnace  may  be 
still  further  modified  by  returning  the  flues  under  the  hearth,  and  thus 
heat  the  ore  from  below.  In  a  long  furnace  this  is  a  doubtful  utility. 
Much  will  depend  on  the  temperature  of  the  furnace  gases. 

Figs.  13,  14,  15,  and  16  show  a  modification  of  the  hand  reverber- 
atory, at  one  time  used  in  Europe  for  roasting  copper  ores. 


Fig.  15.  Fig.  16. 

Figs.  15  and  16. — Modified  hand  reverberatory.     Transverse' sections  through  A-B 

and  C-D  (Fig.  13). 

Mechanical  Reverberatories. — ^Practically  all  the  roasting  at  the 
present  time,  in  preparing  ores  for  treatment  by  wet  methods,  is  done  in 
mechanically  operated  reverberatories.  With  the  exception  of  the 
rabbling  mechanism,  these  furnaces  are  not  essentially  different  from  the 
hand  reverberatories. 

The  mechanical  reverberatories  differ  mostly  from  each  other  in  the 
rabbling  mechanism.  They  may  have  one  long  roasting  hearth,  or 
several  hearths  superimposed  above  one  another. 

The  single  hearth  furnaces  have  the  advantage  of  making  less  dust 
than  the  multiple  hearth  roasters;  while,  on  the  other  hand,  the  multiple 
hearth  furnaces  are  more  compact  and  better  conserve  the  heat. 

In  roasting  ores  for  chemical  treatment,  where  a  more  or  less  com- 
plete roast  is  necessary,  single  hearth  furnaces  have  usually  been  pre- 
ferred, largely  on  account  of  the  low  dust  loss.  Multiple  hearth  furnaces 
have  also  been  largely  used,  and  their  use  is  becoming  more  general. 

In  roasting  pyritic  concentrates  or  heavy  sulphide  ore  down  to  6  or 
8  per  cent,  sulphur,  which  may  be  done  without  fuel,  the  multiple  hearth 
roasters  have  found  more  favor  than  the  single  hearth  furnaces. 
Recently  the  multiple  hearth  furnaces   have  been    modified    by   the 


addition  of  fire-boxes  for  the  different  hearths,  so  that  the  sulphur  can 
be  eliminated  quite  as  completely  as  in  single  hearth  furnaces. 

The  advantages  of  the  multiple  hearth  furnace  is  that  the  heat  from 
the  ore  and  fire-boxes  in  the  lower  hearths,  heats  the  bottom  of  the  upper 
hearths,  and  this  heat  is  very  effectively  applied.  The  disadvantage  is 
that  when  ores  are  roasted  with  external  fuel,  the  volume  of  air  necessary 
for  the  combustion  of  both  fuel  and  the  sulphur  in  the  ore  is  very  large, 
and  the  ore  dropping  from  one  hearth  to  another  against  a  strong  current 
of  ascending  gases,  makes  considerable  dust,  much  of  which  is  lost 
unless  efficient  means  is  provided  for  its  recovery.  This  difficulty  may 
to  some  extent  be  obviated  by  providing  separate  flues  for  the  ore  and 
gases  between  the  hearths,  or  by  exhausting  the  gases  from  one  or 
more  of  the  intermediate  hearths  as  well  as  from  the  upper  hearth. 

In  roasting  heavy  sulphide  material  down  to  6  or  8  per  cent,  sulphur, 
fuel  is  not  ordinarily  necessary,  hence  the  volume  of  air  and  gases  going 
through  the  furnace  is  comparatively  small  and  its  rate  of  passage  com- 
paratively slow,  hence  the  dust  loss  in  dropping  the  ore  from  one  hearth 
to  the  next  need,  not  be  serious  or  excessive.  In  roasting  ores  for  treat- 
ment by  wet  methods,  the  furnace  operates  under  much  higher  tem- 
peratures than  in  roasting  for  smelting,  where  the  sulphur  most  difficult 
to  eliminate,  remains  in  the  ore.  Usually  furnaces,  for  sweet  roasting, 
work  under  a  temperature  of  from  1200  to  1700°  F.  The  ore  itself  may 
not  be  at  these  temperatures,  but  the  reverberatory  chamber  in  which 
the  rabbling  mechanism  operates,  is.  The  temperature  of  the  rever- 
beratory chamber,  at  the  fire-boxes  may  even  exceed  1700°  F.  in  regular 
roasting  work,  and  for  that  reason,  if  anything  goes  wrong  with  the 
rabbling  mechanism,  it  is  necessary  to  at  once  cool  the  furnace  or  the 
top  layer  of  ore  would  take  the  same  temperature  as  the  reverberatory 
gases,  become  sintered,  and  be  unfit  for  extraction  by  hydrometallur- 
gical  processes. 

The  problem,  therefore,  in  mechanically  roasting  ores,  has  been  to 
provide  a  rabbling  mechanism  which  will  work  in  grit  and  dust  and  at  a 
temperature  varying  from  1200  to  1700°  F.,  without  serious  delays  or 
excessive  repairs.  The  different  mechanical  roasting  furnaces  are 
based  fundamentally,  on  the  method  of  overcoming  these  difficulties. 

O'Harra  dragged  the  rabbles  through  a  straight  line  reverberatory 
by  means  of  a  chain  and  track  within  the  reverberatory  chamber.  Brown 
conceived  the  idea  of  placing  a  supplemental  chamber  on  both  sides  of 
the  reverberatory  hearth,  containing  the  tracks  on  which  run  the  rabble 
trucks,  shielded  more  or  less  from  the  direct  heat  and  dust  of  the  rever- 
beratory chamber.  Holthoff-Wethey  placed  the  trucks  entirely  outside 
of  the  furnace  and  devised  a  slot  arrangement  which  opened  and  closed 
automatically  as  the  rabble  moved  along.  Pearce  supported  the  rab- 
bling mechanism  by  a  column  in  the  central  open  space  of  two  concen- 


trie  walls  forming  the  reverberatory  chamber,  the  inside  wall  of  which 
is  slotted  and  the  reverberatory  arch  supported  from  above.  Edwards 
and  Merton  have  a  longitudinal  series  of  rabbles  of  a  diameter  correspond- 
ing to  the  width  of  the  hearth,  projecting  through  the  reverberator^' 
arch,  and  as  the  ore  is  advanced  by  one  of  these  rabbles,  it  is  delivered 
to  the  next,  and  so  on  until  it  issues  from  the  furnace.  The  multiple 
hearth  furnaces  are  mostly  based  on  the  support  and  protection  of  the 
central  column  carrying  the  rabbles  for  the  different  hearths  and  the 
arrangement  for  cooling  and  replacing  either  the  rabble  arms  or  the 

Any  single  hearth  reverberatory  may  be  constructed  with  multiple 
hearths,  but  when  so  modified,  it  is  questionable  whether  they  are  as 
efficient  and  present  the  same  advantages  as  the  regular  circular  multiple 
hearth  furnace  of  the  McDougall  type. 

For  furnaces  where  the  rabbles  are  alternately  within  and  without  the 
roasting  chamber,  as  in  the  O'Harra,  Brown,  and  Holthoff-Wethey,  no 
special  cooling  provision  is  necessary,  since  the  rabbles  never  get  danger- 
ously hot,  and  never,  except  in  case  of  accident,  do  the  rabbles  have  the 
same  temperature  as  the  reverberatory  chamber.  When  the  rabbling 
mechanism  remains  in  the  furnace  indefinitely,  as  in  the  Pearce,  Edwards, 
Merton,  McDougall,  Herreshoff,  and  Wedge,  water  cooling  is  necessary  or 
desirable.  Air  cooling,  for  sweet  roasting,  has  not  given  satisfactory 

There  is  no  great  difference  in  the  cost  of  installation  of  the  various 
mechanical  reverberatories,  and  on  the  same  ore,  for  a  thorough  roast  for 
wet  processes,  there  is  no  great  difference  in  the  cost  of  operation. 

In  roasting  Cripple  Creek  ore,  for  example,  there  are  three  100-ton 
Pearce  roasters  in  one  reduction  works;  six  100-ton  Holthoff-Wethey  at 
another,  and  eight  100-ton  Edwards  at  still  another,  all  in  successful  and 
satisfactory  operation,  roasting  in  all  about  1500  tons  of  ore  daily  and 
having  a  precious  metal  content  of  approximately  $30,000. 

Cost  of  Mechanical  Reverberatories. — The  average  cost  of  a  good 
mechanical  reverberatory  is  about  115  per  square  foot  of  hearth  area  for 
the  smaller  sizes,  and  about  $12  per  square  foot  of  hearth  area  for  the 
larger  sizes,  installed,  ready  to  operate,  but  not  housed. 

Fuel  Required  in  Roasting. — The  fuel  consumption  in  wood  or  coal 
will  usually  be  from  10  to  15  per  cent,  of  the  weight  of  the  ore  roasted, 
and  will  be  more  or  less  independent  of  the  original  sulphur  content.  For 
ore  low  in  sulphur,  considerable  fuel  is  required  to  bring  it  to  the  roasting 
temperature  and  to  maintain  it  at  that  temperature.  For  ore  high  in  sul- 
phur, the  sulphur  itself  develops  considerable  heat,  so  that  extreme  firing 
is  not  necessary  except  to  remove  the  last  few  per  cent,  of  sulphur.  On 
sulphide  concentrates  or  heavy  sulphide  ore,  the  sulphur  content  in  the 
roasted  material  may  be  reduced  to  6  or  8  per  cent,  without  any  fuel,  and 


to  4  and  6  per  cent,  with  only  one  fire-bo-x  at  the  finishing  end  of  the 
furnace.  It  is  in  the  elimination  of  the  last  few  per  cent,  of  the  sulphur 
that  most  of  the  fuel  is  consumed.  But  if  the  fuel  consumption  is  more 
or  less  independent  of  the  sulphur  content  of  the  ore,  the  capacity  of  the 
furnace  is  more  or  less  proportion  to  the  contained  sulphur. 

Hearth  Area  Required  in  Roasting  Various  Ores.— The  capacity  of  a 
roasting  furnace  is  dependent  on  the  amount  of  sulphur  in  the  raw  ore, 
and  on  the  amount  of  sulphur  to  be  eliminated.  For  roasting  ores 
suitable  for  the  hydrometallurgical  processes,  the  hearth  areas  required 
are  approximately  as  follows: 

For  silicious  ore  containing  from  1.5  to  3.5  per  cent,  sulphur  will 
require  from  10  to  15  sq.  ft.  of  hearth  area  per  ton  of  ore  roasted  per 
day;  ores  containing  from  10  to  15  per  cent,  will  require  a  hearth  area 
of  from  25  to  30  sq.  ft.,  and  pyritic  concentrates  which  carry  from  35  to 
45  per  cent,  sulphur  will  require  from  40  to  50  sq.  ft. 

Pyritic  concentrates  and  heavy  sulphide  ores,  carrying  28  per  cent, 
or  more  of  sulphur  are  self  roasting  down  to  6  or  8  per  cent.  After  that 
fuel  has  to  be  used  to  complete  the  roast,  to  make  the  ore  suitable  for 
a  solvent  process. 

The  Brown  Furnace. — The  latest  and  most  approved  type  of  Brown 
furnace  is  shown  in  Figs.  17  and  18,  which  is  a  straight  line  reverberatory, 
in  which  the  ore  is  stirred  and  advanced  by  rabbles  mounted  on  trucks 
attached  to  endless  chains,  one  on  each  side  of  the  furnace,  moving  in 
a  supplemental  chamber.  Brown  was  the  first  to  work  out  a  successful 
method  of  protecting  the  rabbling  mechanism  in  a  straight  line  furnace 
from  the  heat  and  dust  of  the  reverberatory  chamber. 

The  Brown  furnace  has  had  various  modifications,  but  in  its  most  ap- 
proved form  it  is  a  single  hearth  reverberatory,  with  supplemental  chambers 
on  each  side  of  the  roasting  chamber,  to  protect  the  rabbling  mechanism. 
The  supplemental  chambers  are  formed  as  shown  in  Fig.  18  by  a  tile 
projecting  up  from  the  hearth,  above  the  level  of  the  ore,  and  a  corres- 
ponding tile,  forming  part  of  the  reverberatory  arch,  projecting  down 
from  above.  This  construction  forms  a  supplemental  chamber  and  leaves 
a  slot  between  the  roasting  and  supplemental  chambers  just  large  enough 
for  the  rabble  arm  to  pass  through.  The  rabble  arm,  at  the  slot,  is  usu- 
ally made  rather  wide  and  about  an  inch  thick  so  that  the  slot  may  be 
as  narrow  as  possible. 

The  rabbles,  extending  from  one  side  of  the  furnace  to  the  other, 
are  mounted  on  trucks  on  each  side,  and  these  trucks  run  on  tracks  in 
the  supplemental  chambers.  The  trucks  are  attached  to  endless 
chains  which  move  about  sprocket  wheels  at  each  end  of  the  furnace. 
One  pair  of  these  sprockets  is  driven  by  means  of  spurs  and  gears,  which 
in  turn  are  actuated  by  a  counter  shaft  driven  by  belt  and  pulley. 

The  rabbles  in  passing  through  the  furnace  stir  and  advance  the 



I.  .'-i^ 

^  g 





ore,  and  issue  quite  hot;  then  elevated  by  the  sprocket  wheels  at  the  other 
end  of  the  furnace  to  a  track  above,  and  returned  to  the  feed  end  where 
they  are  again  lowered  by  the  driving  sprockets  and  enter  the  reverbera- 
tory  chamber  to  again  complete  the  circuit.  While  returning,  outside 
of  the  reverberatory  chamber,  the  rabbles  are  cooled  enough  so  that  no 
special  cooling  device  is  necessary.  As  each  rabble  enters  the  furnace 
it  takes  with  it  from  the  feed  bin  the  proportionate  amount  of  ore 
required  to  make  the  daily  output.  This  may  be  regulated  by  the  feeding 
device  or  by  the  depth  of  raw  ore  in  the  path  of  the  rabble  as  it  enters 
the  furnace. 

Counterweighted  sheet  iron  doors  at  both  ends  of  the  furnace, 
hinged  at  the  top,  keep  out  the  cold  air;  they  remain  closed  except  when 
lifted  by  the  rabbles  in  passing  in  and  out.  When  the  doors  are  opened, 
even  momentarily,  a  strong  inward  draft  is  likely  to  set  in;  to  avoid  this, 
two  doors  are  sometimes  inserted  at  each  end  a  short  distance  apart  so 
as  to  form  a  neutral  air  chamber;  one  door  being  always  closed  while 
the  other  is  opened  by  the  moving  rabble. 

The  skewbacks  of  the  arch  are  steel  channels  supported  by  the 
buckstaves  and  short  cast  iron  columns,  the  spaces  between  the  columns 
being  3  ft.  6  in.  long  by  12  in.  high.  These  openings  extend  the  entire 
length  of  the  hearth;  they  are  closed  by  sheet-iron  doors,  lined  with 
asbestos.  This  construction  permits  of  ready  access  to  the  hearth  at  any 
point  for  repairs  and  observation. 

From  the  ground  to  the  hearth  the  furnace  may  conveniently  be 
built  of  concrete  or  uncut  stone  and  the  remainder  of  the  furnace  con- 
structed of  brick.  The  furnace  is  bound  together  by  steel  I-beams  which 
also  carry  the  tracks  on  the  top  of  the  furnace. 

The  furnace  is  regularly  made  with  a  roasting  hearth  10  ft.  wide,  and 
in  lengths  varying  from  60  to  200  ft.,  or  even  longer.  The  roasted  ore 
may  be  elevated  to  the  top  of  the  furnace  to  a  sheet  iron  hearth  supported 
by  the  steel  I-beams,  and  there  cooled  and  advanced  by  the  returning 
rabbles,  or  the  roasting  hearth  may  be  somewhat  extended  and  used  as 
a  cooler,  unless  it  is  preferred  to  cool  the  ore  independently  of  the  furnace 

The  frequency  of  stirring  the  ore  may  be  regulated  both  by  the  speed 
of  travel  and  the  number  of  rabbles  on  the  moving  chains. 

The  size  of  the  sprockets  is  governed  by  the  vertical  distance  between 
the  roasting  and  cooling  hearths.  In  driving  the  chain  mechanism  both 
spockets  at  the  drive  end  are  made  tight  to  the  shaft,  while  at  the  other 
end  one  of  the  sprockets  is  tight  to  the  shaft  and  the  other  loose,  so  that 
any  uneven  strain  in  the  chain  is  self  adjusting  and  prevents  the  chain 
from  riding  the  sprockets  and  being  thrown  off. 

The  material  for  a  standard  straight  line  Brown  furnace  10  ft.  wide 
by  140  ft.  long,  actually  erected,  was  as  follows: 


Weight  of  ironwork,  63,000  lb. 
Weight  of  tiles,  35,000  lb. 

The  Brick,  etc.,  required  for  this  furnace  are  as  follows: 

For  Hearth 

46,200  common  red  brick, 

420  partition  tiles, 

14  cu.  yd.  of  sand, 

16,800  lb.  of  ground  fire  clay, 

11  barrels  of  cement, 

126  cu.  yd.  of  stone  for  concrete  work. 

Mortar  for  Hearth;  5  cu.  yd.  of  sand  and  42  bu.  of  lime. 

For  Arch 

1260  common  red  brick, 
1344  skewback  brick. 

For  Five  Fire-Boxes  (one  double  and  three  single  fire-boxes) 
15,000  common  red  brick, 
10,000  common  fire  brick, 
509  arch  tiles  12  in.  long, 
115  arch  tiles  6  in.  long, 
10  fire-door  tiles. 
For  Mortar;  2  1/2  cu.yd.  of  sand  and  25  bu.  of  lime. 

Power. — The  Power  required  to  drive  the  furnace  mechanism,  etc. 
was  supplied  by  a  9  in.  by  12  in.  slide  valve  engine,  which  is  rated  at 
about  25  h.  p.  but  the  engine  was  never  taxed  to  its  capacity  at  any  time. 

Grate  Area. — The  total  area  of  the  grates  in  the  five  fire-boxes  is 
about  63  sq.  ft. 

In  supplying  the  above  material  5  per  cent,  extra  was  allowed  for 
tile  and  skewback  brick,  and  10  per  cent,  on  the  remainder  of  the  items. 

H.  0.  Hof  man'  gives  the  following  data  on  roasting  with  Brown  furnaces 
with  a  hearth  8  ft.  wide  and  135  ft.  long : 

Silicious  ore  with  pyrite,  crushed  to  30  mesh  and  containing  2.5  per 
cent,  sulphur  was  roasted  at  the  rate  of  95  tons  per  24  hours  with  6  cords 
of  wood;  sulphur  in  roasted  ore,  0.3  per  cent.;  roasted  ore  per  square 
foot  of  hearth  area,  131  lb.;  ratio  of  hearth  to  grate  area  32  to  1;  15.83 
tons  of  ore  were  roasted  per  cord  of  wood. 

Silicious  ore  with  pyrite  crushed  to  20  mesh,  containing  2.3  per  cent, 
sulphur,  was  roasted  at  the  rate  of  65  tons  in  24  hours,  with  6  cords  of 
wood;  sulphur  in  roasted  ore,  0.5  per  cent.;  ore  roasted  per  square  foot 
of  hearth  area,  90  lb.;  ratio  of  hearth  to  grate  area, 32  to  1;  10.85  tons  of 
ore  were  roasted  per  cord  of  wood. 

'  "The  MetaUurgy  of  Lead,"  p.  185. 



r  Air  Passage  at  sides  of 
I-BeumB  U>  Cooler  Hcaitb 

Coal  II upper 


Figs.  19  and  20. — Plan  and  section,  Pearce  furnace. 
With  combined  roasting  and  cooling  hearth. 


Silicious  ore  with  pyrite  crushed  to  16  mesh  and  containing  2.5  per 
cent,  sulphur  was  roasted  at  the  rate  of  50  tons  per  24  hours  with  4.5 
tons  of  refuse  slack  coal;  sulphur  in  roasted  ore  0.5  per  cent.;  ore  roasted 
per  square  foot  of  hearth  area,  77.0  lb.;  ratio  of  hearth  to  grate  area 
28.66  to  1;  per  cent,  of  fuel  on  ore,  8.5. 

Copper  matte,  crushed  to  40  mesh  and  containing  20  per  cent,  sulphur 
and  40  per  cent,  copper,  and  15  per  cent,  lead,  was  roasted  at  the  rate 
of  20  tons  per  day;  sulphur  in  roasted  ore,  6.0  per  cent,  the  aim  of  the 
roast  being  to  peroxidize  the  iron  and  convert  as  much  as  possible  of 
the  copper  into  soluble  sulphate;  ore  roasted  per  square  foot  of  hearth 
area,  38  lb.;  ratio  of  hearth  grate  area,  35  to  1;  coal  consumed  3.25  tons; 
per  cent,  of  fuel  on  ore,  16. 

The  Pearce  Furnace. — ^The  Pearce  furnace.  Figs.  19  and  20,  is  of  the 
circular  hearth  type.  The  hearth  is  formed  by  two  concentric  walls, 
usually  about  10  ft.  apart.  The  operating  mechanism  is  at  the  center, 
from  which  the  rabble  arms  radiate. 

The  ore  is  fed  into  the  furnace  from  a  hopper,  to  a  tapering  screw 
located  beneath  the  hearth,  which  distributes  and  raises  the  ore  across  its 
width,  so  that  each  rabble  blade  takes  its  proportionate  share  as  it  comes 
along.  The  rabble  blades  are  attached  to  the  horizontal  pipe  rabble  arms, 
which  in  turn,  are  attached  to  a  rigid  iron  framework  radiating  from  a 
hollow  hub,  at  the  center.  The  hub  revolves  on  ball  bearings,  around 
a  stationary  cast  iron  column. 

The  rabble  arm  is  made  of  two  concentric  pipes,  the  smaller  one  being 
fitted  into  the  larger,  and  having  holes  at  the  end.  These  rabbles  are 
continuously  water  cooled  by  a  gravity  system.  The  water  from  the 
main  is  run  into  a  trough  located  above  the  level  of  the  rabbles  and  re- 
volving with  the  rabbling  mechanism.  It  then  flows  by  gravity  into 
the  inner  concentric  pipe  of  the  rabble  arm  to  the  further  end,  where  it 
is  delivered  to  the  rabble  pipe,  which  is  exposed  to  the  heat  of  the  fur- 
nace, and  flows  back  again  to  the  other  end  of  the  rabble  arm,  from 
whence  it  is  exhausted  into  a  stationary  circular  trough  about  the  hub, 
■near  the  floor.  A  complete  and  continuous  circulation  is  kept  up  in 
this  way.  Air  was  at  first  used  in  the  rabble  arms,  but  for  oxidizing  roast- 
ing, satisfactory  for  wet  methods,  it  had  to  be  abandoned. 

The  inner  circular  wall  of  the  furnace  has  a  continuous  slot  for  the 
passage  of  the  rabble  arms.  It  is  made  reasonably  tight  by  travelling 
.  steel  shields,  counterweighted  so  as  to  press  gently  against  the  walls 
forming  the  slot.  The  upper  part  of  the  wall,  above  the  slot,  rests  on  a 
casting  suspended  from  I-beams  and  cross-beams,  and  braced  by  radial 
struts  and  angle  irons.  The  I-beams  are  supported  by  the  other  wall 
and  central  column,  and  the  cross-beams  by  the  I-beams.  The  outer 
wall  is  18  in.  thick. 

The  rabble  blades,  which  are  made  of  3/8-in.  sheet  steel,  are  graduated 


in  length  and  direction  from  the  inner  to  the  outer  circle,  so  that  the  ore 
on  the  outer  circle,  which  has  to  travel  a  greater  distance,  may  be  at 
the  same  height  and  remain  in  line  with  that  near  the  inner  circle.  In 
other  words,  the  rabbles  are  so  proportioned  that  all  the  ore,  through  a 
cross-section  of  the  furnace,  travels  uniformly  through  the  furnace  and 
is  discharged  at  the  same  time,  notwithstanding  that  the  ore  in  the  outer 
diameter  has  considerably  further  to  travel. 

In  a  furnace  having  an  outside  diameter  of  60  ft.,  the  smallest  blade 
is  3/4X6X3/8  in.,  and  the  largest  blade  is  8X6X3/8  in.,  and  the  inter- 
mediate blades  are  proportioned  to  these  dimensions.  There  are  six 
rabble  arms,  and  each  rabble  arm  has  20  blades,  for  a  hearth  10  ft.  wide. 

The  rabble  arms  are  fastened  to  the  radial  struts  by  a  clamp  which 
permits  the  raising  or  lowering  of  the  rabble  arm,  so  that  the  bed  of  ore 
may  be  kept  level  by  raising  or  lowering  the  further  end  of  the  rabble. 
If  the  rabble  is  not  properly  adjusted,  the  ore  may  pile  up  on  one  side 
of  the  furnace,  instead  of  having  a  uniform  thickness. 

The  depth  of  the  ore  varies  from  2  1/2  to  3  1/2  in.,  depending  on  the 
amount  being  roasted;  2  1/2  in.  is  the  depth  when  roasting  from  80  to 
90  tons  per  day,  and  3  1/2  in.  when  roasting  100  tons  or  more  in  a 
furnace  having  an  external  diameter  of  60  ft.  Such  a  furnace  has  six 
rabble  arms,  which  make  three  complete  revolutions  in  5  minutes,  or 
one  revolution  in  13/4  minutes.  The  ore  is  therefore  rabbled  every 
17  seconds.  The  life  of  the  rabble  arm,  which  is  a  heavy  4-in.  pipe,  is 
one  and  one-half  years.  The  life  of  the  3/8-in.  sheet  steel  blades  is  three 
months  in  roasting  low  sulphur  silicious  ore,  and  from  five  to  six  weeks 
in  roasting  30  to  40  per  cent,  sulphide  ore.  The  blades  are  changed 
without  cooling  the  furnace,  and  without  interfering  with  the  daily 
output  of  roasted  ore.  The  rabble  arms  may  be  changed  by  cooling  the 
furnace  somewhat. 

Two  to  three  per  cent,  sulphur  ore  remains  in  the  furnace  from 
2  1/2  to  3  hours,  and  on  the  cooling  hearth,  1  hour. 

The  height  from  the  top  of  the  ore  to  the  spring  of  the  arch  is  16  in., 
and  from  the  spring  to  the  crown  of  the  arch,  12  in. 

The  Pearce  furnaces,  for  sweet  oxidizing  roasting,  are  usually  built 
with  one  roasting  hearth  above  and  a  cooling  hearth  below. 

The  following  gives  a  summary  of  data  on  the  largest  size  Pearce 
roasters  Figs.  19  and  20,  when  roasting  ores  low  in  sulphur. 

Outside  diameter,  60  ft. 

Width  of  hearth,  10  ft. 

Average  length  of  hearth,  168  ft. 

Mean  diameter  of  hearth,  53  1/2  ft. 

Total  hearth  area,  1689  sq.  ft.     Available  for  roasting  1500  sq.  ft. 

Number  of  fireplaces,  4.     Or  3  fireplaces  and  one  oil  burner. 


Grate  area  of  each  fire  place,  4  ft.  by  4  ft.  8  in. 

Numlier  of  rabble  arms,  6. 

Number  of  blades  on  each  rabble  arm,  20. 

Angle  of  blades  22  1/2  degrees. 

Depth  of  ore,  from  2  to  3  1/2  in. 

Rabbles  make  one  revolution  in  100  seconds. 

Ore  stirred  every  17  seconds. 

Capacity  per  24  hours,  roasting  2.5  to  3.0  per  cent,  sulphur  ore 

down  to  0.5  or  0.75  per  cent.,  100  tons. 
Power  required  to  drive  furnace,  6  h.  p. 
Fuel  required,  10  tons  good  bituminous  coal. 

The  Holthoff-Wethey  Furnace.— The  Holthoff-Wethey  furnace, 
shown  in  Figs.  21  and  22,  is  regularly  constructed  with  a  roasting  hearth 
above  and  a  cooling  hearth  below.  The  roasting  chamber  is  supported 
on  structural  steel,  arranged  so  that  there  is  a  space  between  the  side 
walls  of  the  furnace  and  the  supporting  posts.  In  this  space,  on  both 
sides,  and  attached  to  the  posts  are  the  tracks  for  the  rabble  trucks, 
running  longitudinally  with  the  furnace.  The  rabble  trucks  are  mounted 
on  chains,  which  are  moved,  raised,  and  lowered  by  sprockets  at  both 
ends  of  the  furnace. 

The  power  for  driving  the  mechanism  is  applied  to  the  shaft  at  one 
end  of  the  roaster  on  which  one  pair  of  sprockets  are  mounted.  The 
driving  mechanism,  rabble  trucks,  chains,  and  tracks  are  all  located 
entirely  outside  of  the  roasting  chamber  and  at  all  times  exposed  to  the 
air.  Theslot  through  which  the  rabble  arms  pass  are  opened  and  closed 
automatically  by  tripping  doors,  or  flexible  steel  sheathing,  as  the  rabble 
progresses  through  the  furnace.  One  half  of  the  number  of  rabble  blades 
of  each  rabble  are  set  at  one  angle  and  half  at  an  opposite  angle,  thus 
removing  all  end  thrust. 

The  reverberatory  arch  is  firmly  held  in  place  between  two  h«avy 
I-beams,  suspended  from  above  and  from  beams  resting  on  the  channel 
iron  posts.  Opposite  posts  are  securely  joined  together  and  take  the 
end  thrust  of  the  arch. 

The  ore  is  fed  evenly  into  the  furnace  at  the  drive  end,  and  after 
travelling  the  full  length  of  the  roasting  chamber,  is  dropped  to  the  cool- 
ing hearth  and  again  carried  back  to  the  charging  end,  thus  allowing  the 
ore  the  same  time  to  cool  that  was  required  to  roast  it.  The  cooling 
hearth  may  be  provided  with  water  pipes  or  water  jackets  to  help  cool 
the  ore,  but  this,  on  account  of  the  length  of  the  cooling  hearth,  will 
usually  not  be  necessary. 

The  furnaces  are  usually  built  from  10  to  12  ft.  wide  and  from  100  to 
130  ft.  long.  The  ordinary  dimensions  are  12  X  120  ft.  Such  a  furnace, 
for  roasting  ores  low  in  sulphur,  is  equipped  with  eight  rabbles,  which 







make  a  complete  revolution  in  4  1/2  minutes,  thus  stirring  the  ore  every 
33.5  seconds.  The  upper  sprocket  shaft,  at  the  drive  end,  is  counter- 
weighted  so  as  to  keep  the  rabble  chains  taut. 

The  iron  work  for  a  10  X  100  ft.  Holthoff-Wethey  furnace,  with  roast- 
ing and  cooling  hearths,  weighs  approximately  130,000  lb.  Such  a 
furnace  usually  has  four  fire-boxes  and  a  grate  area  of  60  sq.  ft. 

In  the  erection  of  this  size  furnace  there  were  required: 

68,000  common  brick, 
4,000  fire  brick. 
It  was  driven  by  a  1-5  h.  p.  motor. 


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Fig.  22. — Holthoff-Welhey  furnace.     Transverse  .section. 

This  size  furnace  erected  at  Dalonega,  Georgia,  roasted  concentrates 
containing  over  15  per  cent,  sulphur  down  to  0.15  or  0.20,  all  but  a  trace 
of  which  was  as  soluble  sulphates.  The  capacity  was  35  to  40  tons  per 
day  of  24  hours.  The  ore  contained  10  to  12  per  cent,  moisture  when 
charged  into  the  furnace.  The  fuel  used  was  12  cords  of  wood,  which 
for  35  tons  per  day  would  equal  0.34  cord  per  ton.  At  Colorado  City 
six  Holthoff-Wethey  furnaces,  12  ft.  wide  and  120  ft.  long  have  been  in 
operation  for  many  years  roasting  Cripple  Creek  ores  for  chlorination, 
with  a  capacity  of  100  per  furnace  per  day. 

The  Merton  Furnace.— The  Merton  furnace.  Figs.  23  and  24,  is  a 
rectangular  multiple  hearth  furnace  in  which  the  rabbling  is  done  by 
vertical  shafts  arranged  in  a  line,  and  passing  through  the  respective 
hearths.     To  these  shafts  are  attached  rabble  arms  of  a  radius  equal  to 




one-half  the  width  of  the  furnace.  The  vertical  shafts  are  a  little 
further  apart  than  the  radius  of  the  rabble  arms,  so  that  their  areas  of 
revolution  intersect,  and  the  ore  is  delivered  from  one  rabble  to  the 
next,  through  the  different  hearths. 

Each  shaft  is  set  in  motion  by  a  worm  gear  or  by  spur  and  pinion, 
and  arranged  to  revolve  at  the  desired  speed — usually  from  one  and 
one-half  to  three  times  per  minute.  A  rabble  arm  is  attached  to  each 
shaft  for  each  hearth.  In  the  adjacent  hearths  the  rabbles  on  each  shaft 
are  placed  at  right  angles  to  one  another.  Three  of  the  rabble  arms  in  the 
same  hearth  are  made  to  point  in  the  same  direction,  while  the  fourth, 
or  end  rabble,  is  at  right  angles  to  them. 

The  finishing  hearth  has  a  special  rabble  to  itself  which  is  larger  than 
those  working  on  the  other  hearths;  it  may  be  water  cooled,  but  for  ores 
not  requiring  a  high  finishing  temperature  the  water  cooling  is  often 
dispensed  with. 

The  entire  furnace  is  enclosed  in  1/4-in.  plates,  well  braced  with 
buckstaves  and  rods. 

For  a  standard  type  furnace  the  total  length  is  32  ft.  6  in.  The  main 
body  of  the  furnace  measures  23  ft.  9  in.  long  inside  the  plates,  8  ft.  wide, 
and  the  height,  with  three  hearths,  is  6  ft. 

The  hearths  are  horizontal,  and  are  not  given  any  inclination  what- 
ever. The  height  from  the  hearth  to  the  crown  of  the  arch,  inside 
measurements,  is  16  1/2  in.,  and  9  in.  at  the  sides.  The  thickness  of  the 
crown  of  the  arch  is  4  1/2  in.  Each  of  the  hearths  has  a  door  opposite 
each  shaft.  At  the  end  of  each  hearth  is  a  slot  connecting  with  the 
hearth  below. 

In  roasting,  the  ore  is  fed  in  at  one  end  of  the  upper  hearth,  and 
passed  from  one  rabble  to  the  next  until  it  reaches  the  other  end,  where, 
discharging  through  the  slot,  the  ore  is  delivered  into  the  next  hearth 
below,  and  is  carried  by  another  series  of  rabbles  to  the  opposite  end, 
again  it  is  discharged  to  the  third  hearth  and  travels  along  it  to  the 
finishing  hearth.  Near  the  end  of  the  third  hearth  is  another  slot  in  which 
is  a  gate  permitting  the  control  of  the  discharge  to  what  is  termed  the 
finishing  hearth.     In  this  hearth  the  roast  is  completed. 

The  weight  of  the  iron  work  of  this  furnace  is  about  12  tons.  About 
10,000  ordinary  brick  and  3000  fire  bricks  are  required  in  its  construction. 
About  2  h.p.  is  required  to  drive  the  mechanism. 

The  Edwards  Furnace. — ^In  the  Edwards  furnace.  Figs.  25,  26  and 
27,  the  ore  is  advanced  by  a  series  of  revolving  rabbles  of  relatively 
small  diameter  and  of  intersecting  areas,  so  that  the  ore  from  the  first 
rabble  is  delivered  to  the  next,  and  so  on  through  the  series  until  it 
issues  as  the  roasted  product. 

The  furnace  is  regularly  made  in  two  general  types;  one  having  a 
single  row  of  rabbles,  known  as  the  "Simplex,"  and  the  other  having  a 



double  row  of  rabbles,  known  as  the  "Duplex."  These  two  types  are 
built  either  "Tilting"  or  with  "Fixed  Hearth."  In  the  tilting  con- 
struction there  is  a  double  continuous  girder,  balanced  on  a  center  sup- 
port. The  girder  supports  the  iron  work  of  the  rabbling  mechanism 
and  the  drive  therefor.     The  furnace  is  adjusted  vertically  at  the  end 

Fig.  25. — Edwards  furnace.     Perspective  view. 

s'l'/J-J* — 3'iV,6-  ->^ 

Fig.  26. — Edwards  furnace.     Transverse  section. 

to  give  the  hearth  any  desired  slope.  The  iron  construction  is  encased 
in  brick.  The  fixed  hearth  construction  is  the  same  as  in  standard 
reverberatory  practice. 

The  tilting  furnace  is  a  straight  single-hearth  reverberatory,  63  ft. 
long  by  9  ft.  wide  over  all,  and  58  ft.  by  6  ft.  5  in.  in  the  clear,  which  is 




designed  foi'  sweet  roasting  of- concentrates  or  sulphide  ores.  The  shell 
of  the  furnace  is  a  rectangular  chamber  of  plate-iron  stiffened  with  angle 
iron;  this  is  lined  inside  with  brick.  Good  common  red  brick  are  found 
suitable  for  ordinary  temperatures,  although  at  the  fire  end  fire  bricks 
may  be  used.  10,000  bricks  are  required  for  the  lining  of  this  furnace. 
On  top  of  the  furnace  is  mounted  gearing  for  driving  the  rabbles, 
the  spindles  of  which  extend  down  through  the  cast-iron  boxes,  built 
into  the  brickwork  of  the  reverberatory  arch.  The  whole  furnace  rests 
on  two  pivots,  one  for  each  side.  These  pivots  are  arranged  near 
the  middle  of  the  length  of  the  furnace,  so  that  by  tilting  the  fur- 
nace from  the  horizontal  to  various  angles,  the  discharge  of  the  ore  may 
be  regulated  according  to  the  rate  at  which  it  is  being  roasted.  The 
mechanism  for  tilting  the  furnace  is  simple,  and  can  be  worked  without 
altering  the  usual  speed  of  the  rabbles.  The  weight  of  the  whole  iron 
work  including  the  shell,  bracing,  shafting,  gearing,  rabbles,  etc.,  is  19 
tons.  No  single  part  of  the  furnace  exceeds  2500  lb.  in  weight,  while 
most  of  the  parts  do  not  exceed  225  lb. 

The  bottom  of  the  furnace  is  supported  on  No.  14  corrugated  steel; 
directly  on  this  is  placed  a  thin  layer  of  non-conducting  material,  and  on 
this  is  laid  the  brick  floor. 

There  are  15  rabbles  placed  side  by  side  along  the  length  of  the  furnace, 
the  blades  of  which  nearly  touch  the  hearth.  Each  rabble  has  a  circular 
path,  the  circumference  of  which  almost  touches  the  brickwork  on  either 
side,  and  as  the  distance  between  the  rabbles  is  a  little  greater  than  the 
radius  of  the  circle  described,  each  rabble  works  ore  almost  up  to  the  heel 
of  its  neighbor,  and  as  each  rabble  rotates  in  the  opposite  direction  to  the 
one  next  to  it — ^they  are  alternately  right-  and  left-hand  rabbles — ^the 
ore  is  not  passed  along  too  rapidly  from  one  end  of  the  furnace  to  the 
other,  but  gets  a  thorough  stirring  and  exposure  to  the  air  as  it 
proceeds  on  its  course. 

The  speed  of  the  first  13  rabbles  from  the  feed  end  is  one  revolution 
per  minute,  while  the  fourteenth  has  two,  and  the  fifteenth,  or  discharge 
rabble,  four  revolutions  per  minute.  All  the  rabbles  being  driven  from 
the  same  shafting,  the  change  in  speed  is  arranged  by  alternating  the 
ratio  of  the  bevel  wheels.  The  speed  of  the  fourteenth  and  fifteenth 
rabbles  is  so  arranged  that  while  they  stir  into  the  areas  of  the  adjacent 
rabbles,  they  do  not  come  in  contact  with  them.  The  object  of  this 
increased  speed  is  to  give  the  ore  a  brisker  stirring  in  the  final  stages  of 
the  roast. 

In  order  to  protect  the  rabbles  in  the  hottest  part  of  the  furnace  from 
the  destructive  action  of  the  high  temperature,  water  is  circulated 
through  the  last  three  to  keep  them  cool;  by  this  means  they  are  found 
to  last  for  years. 

There  are  two  different  kinds  of  rabbles  used;  one  is  solid  and  flat- 


footed,  the  front  edge  of  which  is  beveled;  the  other  is  hollow;  on  the 
arm  of  it  cast-iron  shoes  are  fitted.  The  latter  rabble  is  the  one  used 
at  the  fire  end  of  the  furnace.  The  cast-iron  shoes  can  be  slid  on  or  off 
the  rabl)lc  arm  without  lowering  the  heat  of  the  furnace  or  removing  the 
rabble  from  it.  When  roasting  concentrates  or  ores  high  in  sulphur, 
the  first  10  rabbles,  counting  from  the  feed  end,  are  generally  of  the 
solid,  flat-footed  type;  these  pass  through  the  ore  close  to  the  hearth 
and  effectively  stir  and  expose  the  particles  of  ore  so  long  as  they  carry 
a  fair  percentage  of  sulphur.  The  last  five  rabbles  are  provided  with 
cast-iron  shoes,  as  by  the  time  the  ore  reaches  this  part  of  the  furnace  it 
has  lost  most  of  its  sulphur  and  is  less  lively.  These  shoes  pass  through 
and  under  the  ore. 

When  roasting  ores  that  do  not  contain  much  sulphur,  shoes  are 
used  on  all  the  rabbles,  although  water  need  only  be  circulated  through 
those  subjected  to  the  greatest  heat.  The  ordinary  flat-footed  rabble 
arm  is  fastened  to  the  spindle  by  placing  the  end  in  a  socket  and  passing 
a  pin  through  both.  The  water  rabbles  have  a  3  1/8  in.  cast-iron  hollow 
spindle  with  a  flange  at  the  bottom,  which  is  bolted  on  to  a  correspond- 
ing flange  on  the  upper  part  of  the  arm. 

The  ore  is  conveyed  by  an  automatic  feeder  from  the  hopper  into  the 
hearth  at  the  upper  end  of  the  furnace;  after  traveling  to  the  lower  end, 
near  the  fire,  the  ore  is  discharged  down  a  pipe,  located  near  one  of  the 
sides;  the  bottom  of  this  pipe  passes  through  and  works  in  a  case  leading 
to  the  conveyor,  to  prevent  any  escape  of  dust.  This  conveyor  pushes 
the  ore  into  the  pit. 

During  the  roasting  air  is  admitted  through  the  holes,  situated  above 
the  fire  bars.  The  fumes  pass  into  the  main  fine  through  a  short  flue, 
which  is  attached  to  the  furnace.  To  allow  for  the  movement  of  this 
short  flue  when  the  angle  of  the  furnace  is  altered,  the  hole  in  the  main 
flue  through  which  it  passes  is  made  larger  than  actually  required  for  a 
nice  fit;  in  order  to  cover  the  space  left  between  the  short  and  main  flues, 
and  prevent  cold  air  from  passing  into  the  latter,  a  sliding  cover-plate 
moves  in  a  frame,  which  is  bolted  onto  the  brickwork  of  the  main  flue. 

The  power  required  to  work  the  furnace  is  from  2  to  3  h.  p. 

The  Duplex  furnace  is  a  stationary  structure,  designed  for  large 
capacity.  The  principle  of  rabbling  is  similar  to  that  adopted  in  the 
tilting  furnace,  but  consists  of  two  lines  of  rabbles  driven  from  two 
horizontal  line  shafts,  and  the  walls  can  be  built  of  brick  or  concrete. 
The  concrete  can  be  brought  to  within  two  layers  of  brick  to  the  fire 
zone.  The  brackets  carrying  all  the  mechanical  superstructure  are 
firmly  fastened  to  anchor  bolts  in  the  furnace  walls,  and  angle  stays 
and  cross  bars  are  so  arranged  as  to  make  the  entire  superstructure  firm 
and  substantial.     Buckstays  and  tie  rods  hold  the  whole  of  the  arch  and 


brickwork  together  with  straps  on  the  face  of  the  walls  to  give  solidity 
to  the  skewbacks  to  protect  the  arch. 

This  furnace  when  complete  and  ready  for  operating  consists  of 
100,000  lb.  of  iron  and  steel  and  70,000  bricks.  Fire  bricks  are  only 
used  when  in  close  proximity  to  the  fire-box. 

The  hearth  area  of  the  standard  Duplex  type,  112  ft.  long  by  13  ft. 
wide  gives  1456  sq.  ft.  of  working  area,  and  the  outside  measurement 
overall  is  116  ft.  by  16  ft. 

The  usual  fall  of  the  hearth  in  this  furnace  is  1/2  in.  to  1  ft.,  and  the 
mechanism  driving  the  furnace  can  be  so  arranged  to  drive  the  rotating 
rabbles  at  various  speeds  so  that  the  roasting  material  can  be  under  the 
control  of  the  roasterman. 

The  furnace  requires  from  6  to  9  h.  p.  to  operate,  according  to  the 
speed  at  which  the  rabbles  are  driven. 

The  water  required  to  cool  the  rabbles  is  400  gallons  per  minute. 

The  McDougal  Furnace. — The  McDougal  Furnace,^  has  long  been 
used  to  roast  copper  ore  and  concentrates  for  smelting,  and  has  been 
modified  by  Herreshoff,  Evans,  Klepetko,  Wedge,  and  others  to  adapt 
it  to  modern  requirements.  This  furnace,  so  successful  in  imperfectly 
roasting  copper  ores  for  smelting,  has  recently  been  improved  to  adapt 
it  to  the  more  thorough  roasting  required  for  the  hydrometallurgical 

The  McDougal  furnace.  Figs.  28  and  29,  is  essentially  a  multiple  hearth 
upright  cylinder  with  central  shaft  carrying  the  radial  rabble  arms.  The 
hearths  are  horizontal  arches  having  discharge  openings  alternately  at  the 
center  of  one  hearth  and  the  periphery  of  the  next.  The  central  revolving 
shaft  is  provided  with  radial  rabble  arms  for  the  different  hearths,  and 
the  rabble  blades  are  so  arranged  at  an  angle  with  each  arm  that  for  the 
odd  numbered  hearths  they  push  the  ore  toward  the  center,  and  on  the 
even  numbered  hearths  toward  the  periphery.  In  doing  so  the  ore  is 
turned  over  and  over  by  the  rabbles  and  describes  a  spiral  coiirse  around 
the  shaft.  The  ore,  in  its  descent  from  hearth  to  hearth,  describes  a 
zig-zag  course  through  the  furnace  from  top  to  bottom,  passing  alter- 
nately through  the  holes  at  the  center  and  at  the  periphery.  !>; 

The  size  of  the  McDougal  furnaces,  as  largely  used  at  Butte,  is  16  ft. 
in  diameter,  and  18  ft.  3  1/2  in.  high.  It  is  sheathed  with  3/8-in.  boiler 
iron  and  lined  with  a  full  course  of  red  brick.  It  has  six  arched  hearths 
with  a  9-in.  spring  and  3  ft.  apart;  each  hearth  has  two  rabble  arms 
making  one  revolution  per  minute.  Each  furnace  has  two  exhaust 
flues  24  in.  in  diameter  and  12  ft.  apart,  passing  out  of  the  roof,  and 
flues  from  three  furnaces  lead  to  one  main  6  ft.  in  diameter,  having 
openings  along  the  top  and  bottom  for  removing  the  flue  dust. 

'H.  O.  Hofman,  Trans.  Am.  Inst.  Mng.  Eng.,  Vol.  XXXIV;   L.  S.  Austin,  Tras., 
Vol.  XXXVII;  Peters,  "Practice  of  Copper  Smelting." 

HO  A  S  TING  F  UR  XA(  'I'JS 


The  cciitial  shaft  of  tlic  fui'nace  is  driven  from  below.  The  cooling 
water  for  the  rabbles  is  forced  down  to  near  the  bottom  of  the  revolving 
hollow  shaft,  which  is  9  in.  in  diameter,  through  a  3-in.  pipe  and  out  to 
the  ends  of  the  horizontal  rabble  arms  through  1-in.  horizontal  pipes. 
In  its  upward  passage  between  shaft  and  pipe  it  takes  up  the  return 

Fig.  28. — AUis-Chalmers  McDougal  furnace  (fire-box  type). 


Perspective  view  and 

water  from  the  rabble  arms  and  discharges  at  the  top  through  two 
spouts  into  a  secondary  launder.  Shafts  and  arms  are  made  up  of 
flanged  sections  to  permit  of  easy  exchange.  Running  the  overflow 
water  at  80°  F.  20  gallons  of  cooling  water  are  required  per  minute  to 
cool  the  rabbles. 



The  two  rabble  arms  of  a  hearth  have  seven  and  eight  cast-iron 
blades;  these  are  8  in.  long  and  6  in.  wide,  and  5/8  in.  thick;  the  lower 
12  in.  of  the  blade  which  comes  in  contact  with  the  ore  are  chilled.  The 
blades  on  the  top  hearth  last  from  25  to  35  days;  those  on  the  sixth,  from 
6  to  8  months. 

The  six  circular  flat  arches  which  form  the  roof  of  one  hearth  and  the 
floor  of  the  next  above,  require  care  in  construction.  There  must  be  a 
large  central  opening  for  the  main  vertical  shaft  which  carries  the 
rabbles,  and  the  brick  at  the  periphery  must  be  well  anchored.  There 
must  also  be  drop  holes  from  each  hearth  to  the  next,  and  these  are 

K 7—^iA: M 

FiQ.  29. — Elevation  and  section.     McDougall  roaster. 

arranged  alternately  at  the  central  opening  and  at  the  extreme  periphery 
of  the  hearths  and  protected  by  iron  castings.  The  first,  third  and 
fifth  hearths  have  one  drop  hole  at  the  center;  the  second  and  fourth 
have  six  and  the  sixth  has  two  drop  holes  near  the  periphery. 

The  hearths  with  peripheral  openings  are  provided  with  a  central  cast- 
iron  ring,  cut  in  halves.  This  ring  circles  the  shaft,  leaving  an  annular 
clearance  space  of  3  in.  The  brick  of  the  hearth  butts  against  the  ex- 
terior of  this  ring  in  its  entire  circumference,  and  is  keyed  into  a  groove 
in  the  ring.  The  center  drop  hole  is  formed  by  stopping  the  brickwork  of 
the  hearth  so  as  to  leave  an  annular  space  of  16  in.  encircling  the  shaft. 
The  peripheral  drop  holes  are  14  in.  wide  on  the  first  and  fifth  hearths 
and  18  in.  wide  on  the  third  hearth,  where  there  is  a  strong  evolution  of 
sulphur  gas. 

In  roasting  at  the  Washoe  smelter  at  Butte,  the  moist  concentrates 


are  dumped  into  the  feed  hopper  on  top  of  the  furnace,  which  holds  33 
tons.     The  composition  of  the  material  is: 

Moisture,  8.1    percent. 

Cy,  7.42  per  cent. 

SiOj,  21.2   percent. 

1^6,  26 . 1    per  cent. 

S,  33.2   percent. 

"■'2O3)  2.7   percent. 

CaO,  0.3    percent. 

98 .  94  per  cent. 
To  this  is  added  5  per  cent,  of  limestone,  of  which  the  diameter  of 
the  largest  piece  does  not  exceed  1  in. 

The  ore  is  fed  contjnously  into  the  furnace  and  is  spread  on  top  of 
the  hearth  to  the  thickness  of  3  in.  by  the  rabbles.  The  ore  dropping 
down  through  the  holes  in  the  hearths  showers  through  the  ascending 
air,  which  actively  roasts  it,  but  at  the  same  time  this  air  current 
carries  away  the  finer  particles  as  dust,  which  amounts  from  4  to  5 
per  cent,  of  the  ore  charged.  The  gases  escaping  from  the  upper 
hearth  have  a  temperature  of  315°  C;  by  the  time  they  reach  the  flue 
chamber  they  have  cooled  to  117°  C. 

The  appearance  of  the  roasting  at  the  different  hearths  is  as  follows: 
On  the  first  hearth  the  ore  is  dropped  at  the  circumference  and,  containing 
6  to  10  per  cent,  moisture,  is  drying  out,  but  attains  no  visible  heat. 
Entering  the  second  hearth  it  still  looks  dark,  but  shows  a  blue  flame 
by  the  time  it  reaches  the  borders  of  the  hearth,  where  it  is  600°  C.  On 
the  third  hearth  some  sparks  show  where  the  rabble  passes,  together 
with  blue  flame,  and  with  a  flame  temperature  of  900°  C.  On  the  fourth 
hearth  the  sparking  has  ceased,  the  ore  having  attained  an  orange-red 
heat.  In  falling  upon  this  hearth  from  the  one  above,  the  ore  as  it  showers 
down  burns  freely,  hastening  the  roasting  by  this  momentary,  but  thorough 
exposure  to  the  ascending  tir.  On  the  fifth  hearth  the  sulphur  is  elimi- 
nated sufficiently  so  that  the  discharge  temperature  is  less  than  the  enter- 
ing temperature;  that  is,  the  ore  is  brighter  near  the  periphery  than  at 
the  center.  On  this  hearth,  the  maximum  temperature  of  960°  C.  is 
attained.  On  the  sixth  and  final  hearth  the  heat  has  become  uniform, 
but  is  lowered  to  860°  C.  As  the  ore  leaves  the  hearth  it  seems  brighter, 
but  speedily  cools  to  660°  C.  as  it  falls,  smoking  freely,  into  the  hopper. 
Fig.  30  shows  diagramatically  the  progress  of  reactions  and  temperatures. 
The  composition  of  the  escaping  gases  is: 

By  weight        By  volume 
per  cent.  per  cent. 

SOj,  4.95  2.25 

SO,,  1.46  0.53 

0,  19.60  18.45 

N,  75.00  78.77 



Thirty-two  pounds  of  air  is  needed  per  pound  of  sulphur  and  since 
13.3  lb.  of  the  latter  is  burned  off  per  minute,  there  is  needed  in  that  time 
6384  cu.  ft.,  which  passes  up  the  central  hearth  openings  at  the  rate  of 
8.8  ft.  per  second.     A  screen  analysis  of  the  flue  dust  shows: 

On  10  mesh  screen,  9 . 7  per  cent. 

Between  10  and  30,  25 . 3  per  cent. 

Between  30  and  80,  30 . 7  per  cent. 

Passing  80,  33.4  per  cent. 

99 . 1  per  cent. 

The  ore  takes  about  2  hours  and  15  minutes  to  pass  through  the 

In  starting  up  a  furnace,  a  small  fire  of  dry,  soft,  long  flame  wood  is 
started  from  the  three  side  doors  of  the  third  and  fifth  hearths.     A  new 

1st  Hearth 

Snd  Hearth 

8rd  Hearth 

4tb  Hearth 

5th  Hearth 


Percentage  ol  Contained  Sulphur 
20^  40^  60» 























1  ® 







6th  Hearth     . 

«  c  201)  40U  eoo  eoa  looo  c 

Fig.  30. — Progress  of  reactions  and  flame-temperature  in  the  McDougall  roaster. 

furnace  is  brought  to  a  darlc  red  heat  in  3  to  4  days,  an  old 
furnace  requires  only  2  days.  Concentrates  are  then  fed.  After  charg- 
ing for  5  or  6  hours,  it  sometimes  happens  that  the  furnace  cools  down 
too  much,  and  this  makes  it  necessary  to  start  on  the  third  and  fifth 
floors  a  new  fire  for  1.5  to  2  hours;  occasionallyfeeding  of  the  ore  is  stopped 
and  half  a  ton  coal  charged.  Under  normal  conditions  a  furnace  does  its 
best  work  when  the  flue  shows  a  depression  in  water  of  0.3  in.     If  it  is 


less  the  furnace  gets  cool.  The  temperature  may  be  regulated  by  the 
admission  of  air;  closing  the  bottom  doors  drives  up  the  heat,  opening 
the  doors  draws  it  down;  opening  doors  higher  up  checks  the  draught. 
The  rate  of  feed  once  settled  upon  is  usually  not  altered,  and  the  number 
of  revolutions  the  rabbles  make  per  hour  remains  the  same. 

A  section  of  six  furnaces  is  tended  to  in  8-hour  shifts  by  one-third 
foreman,  one  furnaceman,  one  helper,  one-sixth  oiler,  and  one-ninth 
repair  man  and  one  trimmer. 

The  dust  which  collects  in  the  flues  connecting  the  furnaces  forms 
4  to  5  per  cent,  on  the  ore,  is  raked  out  every  day.  The  loss  in  weight, 
including  the  flue  dust,  is  about  20  per  cent. 

A  furnace  treats  under  normal  conditions,  40  tons  of  sulphide  ore, 
with  35  per  cent,  sulphur,  and  10  per  cent,  copper,  or  0.042  tons  per 
square  foot  of  hearth  area,  reducing  the  sulphur  to  7  per  cent.  Roasted 
ore  with  14  per  cent,  copper  treated  in  the  same  manner,  retains  about 
10  per  cent,  sulphur.  The  product  can  of  course  be  varied  with  the  speed 
of  travel  of  the  rabbles,  and  the  sulphur  more  thoroughly  eliminated  by 
the  addition  of  fireplaces. 

The  following  partial  analysis  of  roasted  ore  represents  two  determina- 
tions from  the  average  day  and  night  samples  taken  during  an  experi- 
mental run  of  15  days.  SiOj,  26.9  per  cent.;  Cu,  18.  3  per  cent,  of  which 
9.9  per  cent,  was  present  as  CuO;  Fe,  30.0  per  cent,  of  which  17.9  was 
present  as  FeO;  S,  9.3  per  cent,  of  which  0.81  was  present  as  SO3. 
At  Butte,  in  the  regular  roasting  of  concentrates  the  results  are: 

Amount  roasted  in  24  hours,  40  tons. 

Sulphur  in  raw  concentrates,  35  per  cent. 

Sulphur  in  calcines,  7  per  cent. 

Hearth  area,  952  sq.  ft. 
Concentrates  roasted  per  square  foot  of  hearth  area,     84  pounds. 

Coal,  none 

Cost  of  roasting,  35  cents. 

The  large  sizes  of  the  standard  McDougal  roasters  have  an  outside 
diameter  of  18  ft.  5  in.,  containing  6  hearths,  with  an  enclosed  fire-box 
under  the  sixth  hearth.  The  weight  of  the  entire  iron  work  for  such 
a  furnace  is  about  90,000  lb.  There  are  required  for  its  construction 
about  37,000  red  brick  and  500  fire  brick. 

At  the  Washoe  smelter  of  the  Anaconda  Copper  Mining  Co.,  64 
McDougal  furnaces  are  in  operation;  these  have  a  height  of  18  ft.  3  in., 
and  an  outside  diameter  of  16  ft.  They  are  enclosed  in  boiler  iron  shell 
3/8  in.  thick  and  are  lined  with  a  full  course  of  red  brick,  leaving  an 
inside  hearth  diameter  of  14  ft.  6  in. 

At  Garfield,  Utah,  where  there  are  24  McDougals  in  operation*  the 
ore  and  concentrate  mixtures  were  generally  such  that  it  was  not  neces- 
sary to  roast  below  10  or  11  per  cent,  sulphur.     At  the  same  time  30 

'  R.  R.  Moore,  E.  and  M.  J.,  May  14,  1910. 


per  cent,  of  the  total  charge  was  added  on  the  fifth  hearth,  which 
gave  an  average  of  55  tons  per  day  at  a  cost  of  22  cents  per  ton. 
Other  averages  of  over  50  tons  per  furnace  day  were  maintained  for 
six  months  at  a  cost  of  less  than  25  cents  per  ton.  The  concentrates 
added  on  the  fifth  hearth  of  the  McDougals  were  high  in  copper. 
They  were  added  there  on  account  of  the  fineness  and  tendency  to 
produce  excessive  amounts  of  flue  dust.  At  Garfield  the  McDougals 
produce  about  6  per  cent,  flue  dust.  This  flue  dust  carries  more  silica 
and  sulphur  and  less  copper  than  the  charge.  Notwithstanding  the 
elaborate  system  of  flues  constructed  at  the  Garfield  plant  there  was 
a  stack  loss  of  about  500  lb.  of  copper  per  day  from  these  roasters. 

The  McDougal  furnaces  are  regularly  built  in  the  "Self -roasting" 
and  "Enclosed  Fire-box"  type.  If  a  more  thorough  roast  is  desired 
in  the  self-roasting  type  for  hydrometallurgical  work  than  can  be  obtained 
without  fuel,  satisfactory  results  are  obtained  by  firing  with  oil,  in  which 
case  the  oil  is  injected  into  the  various  hearths,  as  desired.  If  solid 
fuel  is  used,  it  is  desirable  to  use  the  enclosed  fire-box  type.  This  type 
has  two  grates  at  the  bottom,  each  having  an  area  of  29  sq.  ft.  or  a  total 
area  of  58  sq.  ft.  of  grate  surface  to  each  furnace.  In  one  furnace 
of  the  enclosed  fire-box  type,  partly  muffled,  for  roasting  pyritic  ore  con- 
taining 45  per  cent,  sulphur  and  reducing  it  to  an  average  of  2  per  cent,  in 
the  calcines,  there  is  obtained  a  capacity  of  14.4  tons  per  day  of  24  hours. 

The  Herreshoff  Furnace. — In  this  furnace.  Fig.  31,  when  used  in 
roasting  pyrites  for  sulphuric  acid  manufacture,  or  for  the  preliminary 
roasting  of  sulphide  ores  for  metallurgical  treatment,  the  rabbles  are  cooled 
with  air,  through  the  central  column.  This  column  in  a  double  ver- 
tical hollow  shaft.  Attached  to  this  shaft  are  one  or  more  arms  at  each 
hearth,  and  the  replacable  rabbles  are  slipped  on  these  arms.  The  air 
for  cooling  the  rabbles  is  forced  into  the  bottom  of  the  column  and  then 
delivered  through  the  central  part  of  the  shaft,  from  which  it  passes  in 
multiple  at  once  to  all  the  arms.  After  cooling  the  arms  it  returns  to  the 
annular  space  between  the  inner  and  outer  shaft,  and  finally  discharges 
at  the  top  of  the  outer  shaft. 

The  temperature  of  the  iron  in  the  shaft  and  arms  is  kept  above  the 
condensing  point  of  acid  to  prevent  corrosion,  and  at  a  point  where  the 
strength  of  the  metal  is  the  greatest.  The  rabbles  are  made  in  sections. 
There  are  five  sections  on  each  arm  of  from  one  to  five  blades  per  section, 
depending  on  their  position  on  the  arm.  The  sections  can  be  slipped 
on  or  off  the  arms  and  any  blade  can  be  taken  out  of  the  section  and 
replaced  without  disturbing  the  rest.  The  hearths  are  made  of  special 
moulded  arch  fire  brick. 

The  shaft  is  driven  from  the  bottom  by  means  of  a  cast  iron  gear  and 
pinion,  and  makes  one  revolution  in  from  70  to  150  seconds,  depending 
on  the  kind  of  roast.     The  six-hearth  furnace,  15  ft.  9  3/4  in.  diameter, 



requires  about  1  h.  p.  Speed  reductions  are  made  by  gear,  reducing 
worm  gear,  or  sprocket,  as  desired.  A  shear  pin  is  provided  in  the  driving 
mechanism  which  acts  as  a  safty  device  in  case  of  undue  strain.  The 
following  table  gives  dimensions  and  data  for  some  of  the  standard 
size  furnaces. 

Fig.  31. — Herreshoff  furnace. 

Outside  diameter 

Number  of 

Hearth  area 
square  feet 

Weight,  metal 

Weight,  special 
fire  brick 

Pounds,  sulphur 
per  24  hours 

11  ft.  7  1/2  in 

lift.  7  1/2  in.... 
115  ft.  9  3/4  in... 
20  ft 







16,000  lb. 
25,000  lb. 
43,000  lb. 
68,000  lb. 
82,000  lb. 

16,000  lb. 

32,000  lb. 

79,000  lb. 
132,000  lb. 
168,000  lb. 

3,000  to    6,000 

4,500  to  12,000 

8,000  to  21,000 

12,000  to  30,000 

20  ft 

16,000  to  42,000 



in  the  above  tables  the  capacities  given  in  pounds  of  sulphur  per 
24  hours  must  be  used  to  form  a  general  idea  only,  as  the  chemical  com- 
position and  physical  character  of  each  ore,  together  with  the  kind  of 
roast  required,  will  have  to  be  determined  for  each  particular  case. 

The  Wedge  Furnace.— The  Wedge  furnace,  Figs.  32  and  33,  has  for 
many  years  been  successfully  used  in  the  east  for  chloridizing  roasting 

"Jlf^'^'^-^B^v.^'oflifflu)!!'                   1 


JH^:  iiininmniHUp                                    | 


':  ^.Vi" 





-  IP***' 






i,   ,SP 



Fig.  32. — Wedge  furnace. 

copper  ores,  and  its  use  is  rapidly  being  extended  into  the  field  of  oxidizing 
roasting.  The  Wedge  furnace  is  of  the  McDougal  type  and  built  in  var- 
ious diameters,  with  one,  three,  five,  or  seven  hearths,  as  may  be  required. 
The  top  of  the  furnace  is  used  as  a  dryer,  and  a  bottom  hearth,  below  the 
roasting  hearths,  may  be  used  as  a  cooler. 

The  ore  or  concentrate  is  fed  to  the  top  of  the  furnace  at  the  peri- 
phery, and  is  mechanically  fed  across  the  top,  entering  the  center  of  the 


furnace  dry  and  hot.     The  feed  entering  the  furnace  is  so  arranged  that 
the  material  forms  a  lute,  making  the  furnace  gas-tight  at  this  point. 

One  of  the  most  distinctive  features  of  the  furnace  is  the  central 
shaft  whicli  is  hollow,  open  at  the  top  and  bottom,  4  ft.  in  diameter, 
and  is  covered  with  tile  which  are  attached  to  and  revolve  with  the 
shaft.  The  advantage  of  the  large  hollow  shaft  is  that  an  arm  can  be 
changed  easily  and  without  losing  heat  in  the  furnace.  The  arms  are 
held  in  position  by  breech  blocks  placed  upon  the  inside  of  the  central 
shaft;  this  makes  it  possible  for  workmen  to  enter  the  central  shaft  while 
the   heat  is  in  the  furnace,  and  remove  the  breech  block,  when  workmen 


^^^^^^^^^^P^^^-  ^^^^^1 







f..    .  .        _ 



Fig.  33. — Rabble  details,  Wedge  furnace. 

on  the  exterior  of  the  furnace  will  withdraw  the  wornout  arm  through 
one  of  the  doors  and  insert  a  new  arm,  when  the  workman  on  the  inside 
of  the  shaft  will  replace  the  breech  block.  All  parts  in  connection  with 
the  rabbling  mechanism  are  interchangeable. 

The  arms  are  built  for  either  water  cooling  or  air  cooling.  Each  arm 
has  its  own  supply  pipe  and  discharge  pipe,  so  that  it  is  possible  for  the 
furnace  operator  to  know  at  all  times  that  each  arm  is  receiving  its 
proper  supply  of  water. 

The  hearths  are  all  level.  This  is  made  possible  by  building  them 
of  specially  shaped  fire  brick.  These  are  pressed  brick  and  the  arches 
are  laid  up  dry.  The  result  is  that  falling  arches  are  eliminated.  In 
furnaces  which  have  been  in  operation  for  eight  years,  the  arches  are  still 
in  good  condition  without  the  expenditure  of  one  dollar  for  repairs. 

The  weight  of  the  central  shaft  and  arms,  including  the  arms  on  the 
dryer  hearth,  is  carried  on  roller  bearings.  This  reduces  the  power 
required  to  operate  the  furnace  to  a  minimum.  The  indicated  power 
on  a  large  21  ft.  6  in.  diameter  furnace,  with  seven  hearths,  is  less  than 
2h.  p. 

The  furnace  is  built  with  either  a  full  steel  shell  or  skeleton  construc- 
tion, as  may  be  desired.  Common  red  brick  may  be  used  in  the  side  walls. 
The  furnaces  may  be  fired  with  oil,  gas,  or  solid  fuel.     When  solid  fuel 




is  used  the  fire-boxes  are  placed  at  the  sides  of  the  furnace,  and  the  heat 
or  gases  to  the  various  hearths  regulated  by  suitable  dampers. 

The  single  hearth  chloridizing  furnace,  made  for  direct  firing  with 
oil,  is  32  ft.  in  diameter  and  has  a  capacity  of  100  tons  of  pyritic  cinder 
per  day  of  24  hours.  The  fuel  used  is  14  gallons  of  oil  or  210  lb.  of  coal, 
per  ton  of  ore  roasted,  or  approximate  10  per  cent,  of  coal  on  the  weight 
of  ore. 

The  multiple  hearth  mufHe  fired  furnace  has  been  successfully  used 
for  sulphating  roast,  in  which,  on  some  ores,  88  per  cent,  of  the  copper 
was  soluble  in  water.  By  the  use  of  a  weak  acid,  which  can  be  made  at 
low  cost  from  the  escaping  gases,  the  extraction  has  been  increased  to 
98  per  cent. 

The  rabbles  and  rabble  arm,  Fig.  33,  can  easily  and  quickly  be  removed , 
and  replaced. 

The  furnaces  are  built  both  with  open  hearth  or  muffle,  and  of  vary- 
ing diameters  and  number  of  hearths.  The  following  table  gives  the 
essential  figures  for  some  of  the  standard  sizes. 


Diameter  outside 

Number  of 
hearths,  sq.  ft. 

Hearth  area, 
sq.  ft. 

Weight  of      ''  Capacity  in  24  hours, 
metal  parts,  lb. :        tons  of  2000  lb. 


9  ft.  9  in 

9  ft.  9  in 

12  ft 

5                             217 
7                ]              304 
5                1              373 
7                             522 
5             !           75!.'; 












2.5         10.8 
3.5         15.0 
4.3         18.6 


12  ft 

6.0         26  1 


16  ft 

8  3         36  0 


16  ft 



11.6         51.0 


20  ft 



14  3         62  0 


20  ft 

20  0        87  0 



21ft.  6  in 

21  ft.  6  in 

22  ft.  6  in 

5                i            1470 
7                           2058 
3                i              978 

16.9         73.5 
23.6       103.0 
11.2        48.9 


32  ft 





No.  12  furnace  is  designed  more  especially  for  chloridizing  purposes, 
and  the  capacity  shown  above  has  been  demonstrated  in  this  service. 

In  the  capacity  column  the  figures  at  the  left  are  based  on  roasting 
pyrites  containing  50  per  cent,  sulphur,  and  reducing  the  sulphur  to 
2  per  cent,  in  the  roasted  ore. 

The  figures  at  the  right  in  the  capacity  column  are  based  on  smelter 
practice  where  concentrate  containing  35  to  38  per  cent,  sulphur  is 
roasted,  the  sulphur  being  reduced  to  from  7  to  9  per  cent. 

When  oil  is  used  as  fuel  it  can  be  introduced  through  port  holes 
anywhere  at  the  sides  of  the  furnace,  as  in  Fig.  32;  if  coal  is  used,  regular 
fire-boxes  are  necessary,  as  shown  in  Fig.  41 . 

The  Greenawalt  Porous  Hearth. — ^The  porous  hearth.  Fig.  34,  invented 
by  John  E.  Greenawalt,  and  patented  in  1906,  is  applicable  to  any 



furnace.  Some  rather  remarkable  results  have  been  obtained  in  roasting 
ores  by  the  use  of  this  device,  both  as  to  saving  of  fuel  and  capacity  per 
square  foot  of  hearth  area. 

The  essential  principle  involved  is  the  method  of  supplying  sufficient 
air  for  the  ready  oxidation  of  the  incandescent  sulphide  particles  not 
directly  exposed  to  the  oxidizing  atmosphere  of  the  roasting  chamber. 
One  of  the  greatest  objections  to  reverberatory  furnaces  is  that  the  heat 
and  air  cannot  be  most  effectively  applied.     The  top  layer  of  the  ore  is 


Fig.  34. — Greenawalt  porous  hearth  furnace. 

exposed  to  the  highly  oxidizing  atmosphere,  but  that  below  the  surface 
is  in  an  atmosphere  which,  if  not  reducing,  is  certainly  not  highly 

To  obviate  these  difficulties  Greenawalt  conceived  the  idea  of  placing 
the  roasting  ore  on  a  porous  bed,  or  filter,  and  percolate  the  air,  either  up 
or  down,  through  the  ore  mass  and  porous  bed.  In  carrying  out  the 
first  experiments  in  a  hand  reverberatory  furnace,  certain  interesting 
results  were  obtained.  The  air,  in  roasting  heavy  sulphide  ores,  was 
not  found  to  be  of  much  benefit  in  the  early  stages  of  the  roasting,  but 
in  the  later  stages  it  proved  of  the  greatest  advantage.  Another  difficulty 
in  roasting  heavy  sulphide  ores  was  that  the  oxidizing  effects  were  so 
violent  that  the  heat  evolved  sintered  the  charge  so  that  the  sintered 
portion  had  to  be  screened  from  the  roasted  ore  before  it  was  learned 
how  to  properly  regulate  the  draft  or  suction.  The  sintering,  or  agglom- 
eration, was  entirely  due  to  the  air  supply,  so  that  if  the  ore  was  to  be 



roasted  without  agglomeration,  the  air  was  percolated  through  the  ore 
and  porous  bed  with  moderation,  while  if  agglomeration  or  sintering 
was  desired  the  air  was  used  with  considerable  suction  or  pressure, 
depending  upon  whether  down-draft  or  up-draft  was  employed.  Roast- 
ing, with  or  without  sintering,  was  found  to  be  purely  a  matter  of  air 
supply.     The  down-draft  gave  more  uniform  results  than  the  up-draft. 

The  amount  of  air  that  can  be  passed  through  a  bed  of  incandescent 
ore,  on  a  porous  hearth,  is  also  surprising;  as  much  as  14,000  cu.  ft.  of 
air  per  hour  have  been  passed  through  a  hearth  10  ft.  square,  continu- 
ously, without  disturbing  the  ore  particles. 

The  following  results  were  obtained  on  tests  made  by  the  New  Jersey 
Zinc  Co.  with  the  demonstration  furnace  at  Denver.  The  furnace  has 
a  hearth  area  100  sq.  ft.  It  is  arranged  for  hand  rabbling  and  fired 
with  coal.  The  tests  were  made  under  the  supervision  of  W.  C.  Wetheril, 
consulting  engineer  and  metallurgist  of  the  Empire  Zinc  Co.,  and  Wm. 
H.  Faul,  assistant  engineer.  The  chemical  determinations  were  made 
by  the  company's  chemist,  E.  M.  Johnson. 



11.00  a.  m.     Charged  raw  ore 

1.00  p.  m. — second  hour 

3.00  p.  m. — fourth  hour 

5.00  p.  m. — sixth  hour 

7.00  p.  m. — eighth  hour 

9.00  p.  m. — tenth  hour 

11.00  p.  m. — twelfth  hour 

1.00  a.  m. — fourteenth  hour 

3.00  a.  m. — sixteenth  hour 

5.00  a.  m. — eighteenth  hour 

7.00  a.  m. — twentieth  hour 

9.00  a.  m. — twenty-second  hour 

11.00  a.  m. — twenty-fourth  hour 

Roasted  ore,  final  average  of  entire  charge. 

Crude  concentrates. . . . 
Roasted  concentrates 

Weight  of  charge,  raw,  2000  lb. 
AVeight  of  charge,  roasted,  1690  lb. 
Shrinkage,  310  lb. 
Coal  consumed  (slack),  3052  lb. 

Total  S     I   Sol.  H2O 


Insoluble  S 







































































9.45  a.  m.     Charged  raw  ore 

10.45  a.  in. — first  hour 

11.45  a.  m. — second  hour 

12.45  p.  m. — third  hour 

1.45  p.  m. — fourth  hour 

2.45  p.  m. — fifth  hour 

3.45  p.  m. — eighth  hour 

4.45  p.  m. — seventh  hour 

5.45  p.  m. — eighth  hour 

6.45  p.  m. — ninth  hour 

7.45  p.  m. — tenth  hour 

8.45  p.  m. — eleventh  hour 

9.45  p.  m. — twelfth  hour 

Roasted  ore,  final  average  of  entire  charge. 

























Crude  concentrates. . . 
Roasted  concentrates. 


Furnace  charge  raw  ore,  weight  1800  lb. 
Furnace  charge  roasted  ore,  weight  1551  lb. 
Shrinkage,  249  lb. 
Coal  (alack)  consumed,  1160  lb. 





In  these  comparative  tests,  alternate  charges  of  ore  were  roasted 
without  and  with  air  passing  through  the  hearth,  other  conditions 
remaining  the  same. 

From  these  tests  it  was  concluded  that,  with  the  porous  hearth  the 
capacity  of  the  furnace  was  increased  from  two  and  one-half  to  three 
times;  that  the  amount  of  fuel  required  was  only  35  per  cent,  of  the 
amount  required  without  the  air,  and  that  the  total  cost  of  roasting  was 
reduced  by  65  per  cent. 

Two  mechanical  furnaces,  10  ft.  wide  and  100  ft.  long  were  then 
erected  to  roast  this  and  similar  material,  but  the  lead  and  other  ingre- 
dients of  the  ore  formed  a  smooth  crust  due  to  the  friction  of  the  rabbles 
with  the  stationary  ore  on  the  hearth,  which  soon  became  impervious 
to  the  air.  To  what  extent  such  a  crust  would  form  in  roasting  other 
ores  is  somewhat  questionable.  In  silicious  ores,  like  those  of  Cripple 
Creek,  no  crust  of  any  kind  is  discernible  between  the  moving  and 
stationary  ore  of  the  hearth,  even  after  furnaces  have  been  in  operation 
for  years.  The  stationary  ore  is  as  permeable  as  the  moving  ore.  Under 
such  conditions  no  crust  difficulties  should  arise,  and  the  capacity  of 
the  furnace  should  be  enormously  increased,  and  the  cost  of  roasting 
materially  diminished. 






Bruckner  Furnace. — ^The  Bruckner  furnace  was  introduced  in  Colo- 
rado in  1867,  and  since  that  time  has  been  more  or  less  generally  used 
bgth  in   oxidizing  and  chloridizing  roasting. 

The  furnace,  as  shown  in  Figs.  35  and  36,  consists  of  a  cylindrical 
shell  of  steel  plate,  with  circular  openings  at  each  end.  Two  circular 
tracks  are  fastened  around  the  cylinder,  at  equal  distances  from  the  ends. 
With  these  tracks  the  cylinder  rests  on  four  riding  wheels  or  rollers, 
which  are  mounted  on  strong  cast  iron  frames.  Two  of  the  rollers  have 
double  flanges  to  keep  the  cylinder  in  proper  alignment. 

A  revolving  motion  is  given  the  cylinder  by  means  of  cogs  and 
pinions,  so  adjusted  that  the  proper  reduction  in  speed  is  made  between 
the  driving  pulley  and  the  revolving  cylinder.  A  large  peripheral  cog 
is  fastened  to  the  cylinder,  which  meshes  into  a  pinion  of  suitable  size. 
Fig.  36  shows  the  mechanical  details  of  the  cylinder  and  driving  mech- 
anism for  a  furnace  8  1/2  ft.  diam.  by  18  1/2  ft.  long,  and  having  a 
capacity  of  8  to  11  tons  of  ore  at  a  charge. 

A  fire-box  is  placed  at  one  end  of  the  cylinder,  the  throat  of  which 
corresponds  with  the  central  end  opening;  while  a  similar  opening  at  the 
other  end  communicates  with  the  stationary  flue  or  dust  chamber.  The 
fire-box  for  the  furnace  is  built  of  brick,  and  in  some  instances  it  is 
mounted  upon  wheels  so  that  it  may  be  removed  to  facilitate  the  relin- 
ing  of  the  shell.  When  a  complete  roast  is  desired,  a  stationary  fire-box 
is  to  be  preferred  over  a  removable  one. 

There  are  four  doors  in  the  cylinder;  two  of  these  are  placed  diamet- 
rically opposite  the  other  two.  The  doors  serve  for  charging  and  discharg- 
ing the  ore.  A  hopper  is  placed  above  the  furnace  large  enough  to  hold 
a  charge  of  the  material  to  be  roasted;  this  hopper  has  two  outlet  spouts, 
each  provided  with  a  slide,  which  corresponds  with  the  furnace  doors. 

The  shell,  as  well  as  the  ends,  are  lined  with  brick.  Provision  is 
made  in  the  driving  mechanism  to  regulate  the  speed  from  one  revolution 
per  minute  to  one  revolution  in  3  minutes. 

The  ends  of  the  cylinder  are  sometimes  contracted  to  facilitate 
discharging  of  the  ore,  but  this  is  not  necessary,  except  for  large  furnaces. 

These  furnaces  are  usually  built  of  various  standard  sizes;  the  di- 
mensions, weights,  and  capacities,  are  approximately  as  follows: 

Size  of  furnace 

6  X12 

7  X18 

8  1/2X18  1/2 
8  1/2X28 

Weight  of  iron  work 

17,800  lb. 
30,000  lb. 
52,000  lb. 
69,000  lb. 

fire  brick 

GommoQ  brick 

Capacity,  per 
charge,  tons 






6  to    8 



8  to  11 



15  to  25 



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mg^ga^kJA   \ 



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In  the  operation  of  the  furnace  two  of  the  doors  in  the  cylinder  are 
opened  and  brought  directly  under  the  hopper  spouts.  The  slides  are 
then  withdrawn  and  the  charge  allowed  to  run  in.  The  furnace  is  then 
given  a  slight  turn  to  bring  it  in  a  convenient  position  to  close  the  doors, 
which  are  then  fastened  as  tight  as  possible. 

The  cylinder  is  then  revolved  and  a  strong  fire  maintained,  until 
the  sulphur  is  well  ignited.  If  the  charge  being  roasted  is  heavy  sul- 
phide ore  or  pyritic  concentrates,  the  firing  is  discontinued  after  the 
sulphur  is  thoroughly  ignited.  The  heat  developed  by  the  burning 
sulphur  is  considerable,  and  if  augumented  by  external  firing,  caking 
and  balling  of  the  ore  would  take  place. 

Usually  two  charges  of  ore  are  roasted  in  24  hours.  If  the  ore  is  a 
heavy  sulphide,  the  furnace  can  be  run  from  three  to  four  hours  with- 
out firing.  If  the  ore  is  being  chloridized,  salt  is  then  added,  and  the 
roasting  continued  with  a  moderate  fire.  The  temperature  may  be 
observed,  and  scoop  samples  taken,  from  a  hole  back  of  the  flue.  When 
the  roasting  has  fairly  well  progressed,  the  charge  in  the  furnace  loses  its 
tendency  to  run  like  a  liquid,  and  assumes  an  inclined  position,  up  to 
45  degrees.  This  unequally  distributed  weight  acts  against  the  direction 
of  the  motion,  and  if  the  clutch  is  thrown  out  in  order  to  stop  the  furnace, 
this  weight  will  pull  the  furnace  back  nearly  a  quarter  of  a  turn.  If  the 
ore  is  given  a  chloridizing  roast,  the  salt  is  introduced  through  two  of  the 
doors  and  well  scattered  over  the  charge. 

When  the  roasting  is  finished  two  cars  are  pushed  under  the  furnace, 
one  for  each  door.  All  four  doors  are  then  opened  and  the  furnace 
again  revolved.  The  receiving  cars  are  made  narrow  so  that  no  ore 
is  dropped  beyond  the  confines  of  the  car,  because  the  ore  discharges 
through  the  doors  over  a  large  arc. 

The  revolving  motion  of  the  Bruckner  furnace  should  be  slow.  The 
ore,  in  any  case,  is  continuously  changing  its  position  and  exposing  new 
surfaces.  There  is  no  advantage  in  moving  the  ore  more  rapidly  than 
the  ore  at  the  surface  can  be  oxidized.  One  revolution  in  two  or  three 
minutes  should  be  ample. 

In  the  larger  sized  Bruckner  furnaces,  the  ore  is  exposed  to  rather 
an  uneven  heat.  That  near  the  fire-box  is  at  a  considerably  higher  tem- 
perature than  the  ore  at  the  further  end,  from  18  to  28  ft.  distant.  To 
overcome  this  defect,  O.  Hofmann  modified  the  Bruckner  furnace  by 
placing  a  fire-box  at  both  ends,  and  arranged  so  that  the  fire-boxes  could 
be  used  intermittently,  by  damper  arrangements  connecting  with  the 
flue,  or  dust  chamber. 

The  cylinders  are  made  to  revolve  slowly;  the  smaller  ones  by  apply- 
ing power  to  a  shaft  carrying  the  friction  rollers,  the  larger  ones  by  a 
pinion  which  engages  a  spur  gear  surrounding  the  cylinder. 

Howell-White  Furnace.— The  Howell- White  furnace.  Figs.  37  and  38, 



consists  essentially  of  a  long  telescope-shaped  cast-iron  cylinder,  made 
in  sections  with  cast  flanges,  carefully  fitted,  and  bolted  together.  The 
cylinder  is  supported  on  a  system  of  rings  resting  on  friction  wheels,  and 
guided  in  a  central  position  by  rollers  in  upright  frames,  and  revolved  by 
friction  of  the  wheels  operated  by  gears  and  pulleys.  The  flame  passes 
through  the  revolving  cylinder  from  a  fire-box  at  one  end  to  the  flue  at  the 

That  one-third  portion  of  the  cylinder  nearest  the  fire  has  a  larger 
external  diameter  than  the  part  next  the  flue,  but  it  is  lined  with  fire 
brick  to  make  its  internal  diameter  the  same  as  that  of  the  smaller  part, 
which,  although  unlined,  stands  the  heat  very  well.  In  some  instances 
the  entire  cylinder  is  lined,  and  this  is  probably  the  best  when  the  ore  is 
given  a  thorough  oxidizing  roast. 

Fig.  37. — Howel-White  furnace. 

Projecting  fire  brick,  arranged  spirally  in  the  brick-lined  portion, 
assist  in  oxidation  by  raising  and  showering  the  ore  through  the  flame, 
which  passes  directly  through  the  cylinder.  When  the  feed  end  of  the 
furnace  is  not  lined  with  brick  cast-iron  projections  are  provided  for  the 
same  purpose.  These  projections  make  large  quantities  of  dust,  some- 
times from  30  to  50  per  cent,  of  the  ore  charged,  and  for  that  reason  are 
frequently  omitted. 

The  cylinder  is  inclined  slightly  toward  the  discharge  end  in  order 
to  advance  the  ore  gradually  against  a  constantly  increasing  temperature. 
The  furnace  is  fed  at  the  upper  end  with  dry  ore  by  means  of  a  screw 
feeder,  and  then  makes  its  way  automatically  toward  the  lower  end  of  the 
furnace,  where  it  passes  out,  dropping  between  the  end  of  the  cylinder 
and  the  fire-box,  into  the  vault. 

On  account  of  the  excessive  dust  from  the  furnace,  an  auxiliary 
fire-box  is  usually  placed  at  the  flue  end  of  the  furnace  for  roasting  the 
flue  dust  as  it  passes,  suspended  in  the  air,  into  the  dust  chamber.  If 
the  spiral  projections  are  omitted  from  the  cylinder,  the  auxiliary  fire 
to  roast  the  dust  may  also  be  omitted.  The  ore  as  fed  into  the  furnace 
has  its  sulphur  fairly  well  eliminated  before  it  comes  against  the  more  or 
less  direct  flame  from  the  fire. 



The  furnace  requires  very  little  repair,  and  very  little  power  to  run 
it,  while  its  capacity  is  quite  large.  The  revolving  speed  of  the  furnace 
is  adjustable;  and  may  vary  from  three  to  eight  revolutions  per  minute, 
the  larger  furnaces  revolving  more  slowly  than  the  smaller  ones. 

Fig.  37  shows  a  longitudinal  elevation  of  the  furnace  and  Fig.  38,  a 
section  through  the  discharge  end. 

-13-10^^ 1 ; > 

Truck  to  5  *  ^ 

WoodYarJ  /        Hrlck  Pftvios 



Wln-'tliiig  Flour 

Fig.  38. — Howel- White  furnace.     Discharge  end  and  ore  vault. 

The  following  table  gives  the  standard  sizes  of  the  Howell-White 
furnace,  with  weights,  material,  capacities,  etc. 





Capacity  in 

15  to  20 
30  to  50 
50  to  60 
60  to  70 

Weight  of  iron 
work,  lb. 


Fire  brick 

Common  brick 







This  furnace  has  been  largely  used  for  chloridizing  roasting;  when 
so  used  the  salt  must  be  fed  in  with  the  ore.  A  portion  of  the  salt  may 
be  added  as  the  ore  drops  into  the  vault,  after  passing  through  the 

Muffle  Furnaces.— All  of  the  well-known  designs  of  standard  rever- 
beratories  may  be  modified  to  mufHe  furnaces  by  placing  flues  under  the 
hearth,  and  by  having  two  concentric  arches  over  the  reverberatory 

A  modification  of  the  hand  reverberatory  for  muffle  roasting,  in 
chloridizing  pyritic  cinders,  is  shown  in  Fig.  49,  page  263. 

Section  D-D 
Fig.  40. — Edwards  muffle  furnace.     Transverse  section. 

A  modification,  of  the  Brown  furnace  for  muffle  roasting  is  shown  in 
Fig.  39;  for  the  Edwards,  in  Fig.  40;  for  the  Wedge,  in  Fig.  41,  and  for 
the  Allis-Chalmers  McDougal,  in  Fig.  42.  These  illustrations  fairly 
represent  the  modifications  required  for  the  straight  line  chain-driven 
roasters;  the  straight  line  circular  rabbled  roasters,  and  the  multiple 
hearth  furnaces  with  central  revolving  shaft. 

Muffle  furnaces  are  not  as  economical  as  reverberatories,  but  for 
certain  work  present  recognized  advantages.  The  heat  can  be  regulated 
more  uniformly  and  the  amount  of  dust  is  minimized.     In  chloridizing 



roasting,  muffle  furnaces  present  the  decided  advantage  that  the  volatile 
metals  in  the  fumes  can  be  more  readily  condensed  and  recovered.  In 
reverberatories,  or  other  direct-fired  furnaces,  the  gases  from  the  fires  are 
mixed  with  those  from  the  ore,  so  that  the  volume  of  gases  passing 
through  the  condensers  for  the  recovery  of  the  volatilized  values  is  quite 
large,  and  by  far  the  largest  amount  of  air  in  such  cases  is  due  to  the 
combustion  of  the  fuel.  In  chloridizing  roasting  only  that  part  of  the 
furnace  containing  the  ore  after  the  salt  is  added  need  be  muffled. 

A  combination  of  reverberatory  and  muffle  can  frequently  be  used  to 
advantage  to  avoid  excessive  dust  loss,  and  where  it  is  intended  to  make 

Fig.  41. — Wedge  muffle  furnace. 

sulphuric  acid  to  leach  the  copper  from  the  roasted  ore.  Fig.  42  shows 
such  a  furnace,  in  which  the  three  lower  hearths  are  muffled  with  special 
tile  to  keep  the  products  of  combustion  from  the  fire-boxes  separate 
from  the  sulphur  dioxide  gases;  the  furnace  is  equipped  with  uptakes  for 
conveying  the  gases,  and  a  separate  stack  is  arranged  with  connections 
to  the  muffles  for  taking  away  the  gases  of  combustion.  If  sulphuric 
acid  is  to  be  made,  furnaces  may  be  muffled  so  as  to  get  the  maximum 
sulphur  dioxide  content  in  the  furnace  gases  for  the  manufacture  of  the 
acid,  and  the  remaining  sulphur  in  the  ore  can  then  be  eliminated  by 
a  continuation  of  the  same  hearth  into  a  reverberatory  chamber  in 
which  the  fumes  from  the  ore  and  from  the  fuel  commingle  and  pass 
out  together,  and  separately  from  the  sulphur  gases  from  the  muffle. 

no  A  H  TI  A'G  F  UR  NA  CES 


^s^^iO.ll  -'.  «\^-\^-.^'I»^;5  ■  .^-.^  ;  .„  isS^^^^Sl^ 

Fig.  -12. — AUis-Clialmcrs  McDougal  muffle  furnace. 


Fig.  41  shows  a  muffle  furnace  now  largely  used  in  the  chloridizing  roast- 
ing of  cupriferous  pyritic  cinders  for  the  extraction  of  the  copper.  This 
type  of  furnace  is  replacing  the  hand  muffle  furnaces  used  for  that 

Ore  Coolers. — After  roasting,  it  is  necessary  to  cool  the  ore  before 
charging  it  into  vats  or  agitators  for  chemical  treatment.  The  cooling 
in  almost  all  of  the  furnaces  is  done  in  combination  with  the  roasting  by 
a  mechanical  device  similar  to  the  rabbling  mechanism  of  the  furnace. 
Usually,  as  in  the  case  of  the  Pearce,  Holthoff-Wethey,  and  Wedge 
furnaces,  by  dropping  the  ore  to  a  lower  open  hearth  and  continuing  the 
rabbling  as  in  the  roasting.  Frequently  the  cooling  hearth  is  made  of 
water  j  ackets,  or  pipes,  through  which  cold  water  is  continually  circulated. 
When  the  design  does  not  conveniently  lend  itself  to  the  use  of  a  lower 
hearth,  the  ore  is  cooled  by  a  rabbling  device  similar  to  the  roasting 
mechanism,  and  either  connected  with  it,  or  as  a  separate  apparatus. 
A  push  conveyor,  with  water  jacketed  bottom,  arranged  to  deliver  the 
cooled  ore  to  the  leaching  plant,  makes  a  good  cooler.  In  Figs.  20  and 
22  the  lower  hearth  of  the  furnace  is  used  to  cool  the  ore. 

Dust. — Dust  presents  one  of  the  most  serious  problems  in  roasting.  ' 
If  the  ore  fed  into  a  furnace  contains  an  excessive  amount  of  dust,  the 
capacity  is  reduced.  In  one  of  the  Cripple  Creek  mills,  when  the  ore  in 
the  supply  bins  became  low  and  contained  an  excessive  accumulation  of 
dust,  the  capacity  of  the  furnaces  was  reduced  from  100  to  90  tons  per 
day,  but  this  decreased  capacity  was  of  short  duration,  as  the  bins  soon 
became  empty  and  a  fresh  supply  of  ore  had  to  be  provided.  Ore 
crushed  exceedingly  fine  cannot  be  roasted  at  the  same  rate  as  ore 
crushed  from  8  to  20  mesh. 

All  furnaces  make  more  or  less  dust  in  roasting,  which  if  not  recovered, 
represents  a  serious  loss;  and  if  recovered,  an  additional  expense.  A 
dust  chamber  is  usually  a  necessary  adjunct  to  a  roasting  furnace,  and 
must  be  taken  into  account  in  a  roasting  installation. 

For  straight  line  reverberatory  furnaces  the  dust  recovered  in  suitable 
dust  chambers  varies  from  1.5  to  3.5  per  cent,  of  the  ore  roasted.  Much 
depends  on  the  nature  of  the  ore  and  the  speed  of  travel  of  the  rever- 
beratory gases.  In  multiple  hearth  furnaces  the  recovery  in  the  dust 
chambers  will  usually  be  from  3.5  to  5  per  cent,  in  careful  work.  In  re- 
volving furnaces,  the  dust  recovery  in  the  flues  or  chamber  is  quite  large. 

M.  W.  von  Bernewitz  gives  as  an  example'  a  battery  of  six  roasting 
furnaces  with  a  capacity  of  20  tons  each  of  Kalgoorlie  ore  daily,  that 
has  been  crushed  to  25-mesh  screen.  These  furnaces  are  connected  with 
a  flue  100X7X7  ft.  and  a  stack  100X6  ft.  with  a  1/2-in.  draft.  Twenty 
tons  of  flue  dust  are  collected  monthly,  which  amounts  to  about  3.5  per 
cent,  of  the  ore  roasted. 

ȣ.  and  M.  J.,  Feb.  26,  1910. 


As  a  rule  the  dust  collected  from  the  dust  chambers  is  not  well 
roasted.  This  is  largely  due  to  the  fact  that  the  partially  heated  ore  dusts 
more  than  after  it  has  been  heated  to  incandescence,  but  more  largely 
due  to  the  rapid  travel  of  the  fuel  and  furnace  gases  over  the  partially 
heated  ore  at  the  rear  of  the  furnace.  Roasting  furnaces,  as  usually 
built,  exhaust  all  the  gases  from  one  opening  at  the  rear  of  the  furnace, 
so  that  as  each  fire-box  discharges  its  gases  into  the  roasting  chamber, 
the  draft,  or  speed  of  travel  of  the  gases  against  the  roasting  ore  must 
be  proportionately  increased  with  each  addition.  At  the  rear  of  the 
furnace  all  the  fire-boxes  have  poured  their  gases  into  the  roasting 
chamber,  and  as  the  ore  is  rabbled,  the  rapid  movement  of  the  gases 
over  it,  whip  a  large  proportion  of  the  finer  unroasted  material  into  the 
dust  chamber. 

There  is  no  reliable  data  available  on  the  dust  loss  through  the  stacks. 
It  is  probably  more  than  usually  supposed.  In  one  of  the  Cripple  Creek 
mills  there  was  at  one  time  an  unaccountable  loss  of  values  amounting 
to  1  per  cent.  It  was  thought  that  much  of  this  was  due  to  dust  losses 
in  the  furnace  stack. 



Roasting  of  Cripple  Creek  Ores. — Most  of  the  ores  in  Cripple  Creek 
are  roasted  before  chemical  treatment.  The  amount  roasted  per  day- 
is  about  1500  tons.  The  ore  after  roasting  is  treated  both  by  the  chlorina- 
tion  and  cyanide  processes.  All  the  furnaces  now  in  use  are  designed  to 
roast  100  tons  of  ore  daily.  Of  the  furnaces  in  regular  operation,  there 
are'  three  Pearce  furnaces;  six  Holthoff-Wethey;  eight  Edwards,  and 
one^Holthoff  revolving  hearth.  All  of  these  furnaces  are  giving  satis- 
factory results. 

The  average  chemical  composition  of  Cripple  Creek  ores  which  are 
roasted,  is  about  as  follows: 

SiOj,  60  to      70  per  cent.  S,  1 . 5  to     3 . 0    per  cent. 

AI2O3,  18  to      22  per  cent.  Pb,  trace         to     0. 10  per  cent. 

CaO,  1 . 5  to    2.5  per  cent.  Zn,  trace        to     0.10  per  cent. 

MgO,  0 . 1  to  0 . 25  per  cent.  Mn,  trace        to     0.15  per  cent. 

BaO,  trace  to  0 .  25  per  cent.  K  and  Na,  5 . 0  to  10.0    per  cent. 

Fe,  3 . 0  to  7.0    per  cent.  Te,  trace. 

Cu,  trace. 

The  ore  as  it  comes  from  the  mines  is  crushed  through  rock  breakers 
and  rolls  to  pass  through  8-  to  16-mesh  screens.  The  oxidized  ore,  as  it 
comes  from  the  mines,  contains  as  low  as  0.25  per  cent,  sulphur;  the  un- 
oxidized,  from  the  deeper  levels,  contains  as  much  as  5  per  cent.;  the 
average  for  the  different  mills  ranges  between  1.5  and  2.75,  depending 
on  the  condition  of  the  mines  from  which  most  of  the  ore  comes. 

Ore  low  in  sulphur — from  1.0  to  2.0  per  cent. — is  roasted  until  it 
contains  between  0.25  and  0.60  per  cent.;  that  carrying  from  2.25  to  3.25 
per  cent,  to  between  0.50  and  0.85  per  cent.  All  of  the  mills  are  at  pres- 
ent equipped  with  furnaces  which  regularly  roast  100  tons  of  ore  con- 
taining from  1.5  to  2.5  per  cent,  sulphur;  when  the  sulphur  exceeds  2.5 
per  cent,  the  capacity  is  likely  to  be  reduced  to  90  tons,  while  on  the 
contrary,  if  the  sulphur  is  about  1.5  per  cent,  or  less,  the  capacity  may 
exceed  120  tons  per  day  of  24  hours.  The  hearth  area  of  these  furnaces 
is  between  1200  and  1500  sq.  ft.;  the  grate  area  about  75  sq.  ft. 

Many  of  the  furnaces  are  fired  with  producer  gas.  The  ore  coolers 
are  usually  located  directly  under  the  roasting  hearths  and  form  part  of 
the  furnace,  so  that  the  ore  on  issuing  from  the  roasting  hearth  drops  to 
the  cooling  hearth,  to  be  turned  over,  cooled,  and  advanced  by  the  cool- 
ing hearth  rabbles. 



The  consumption  offuel  varies  fronil0tol5tonsofWestern  bituminous 
coal  per  100  tons  of  ore  roasted.  Lignite  is  used  to  some  extent,  but  is 
not  as  effective  as  the  longer  flame  bituminous.  Oil  and  residuum  are 
also  used,  but  in  connection  with  coal.  If  they  are  used  together  it 
takes  5  tons  of  coal  and  150  gallons  of  oil  per  100  tons  of  ore. 

The  ore,  in  roasting,  is  given  an  initial  temperature  at  the  first  fire-box 
of  1200  to  1300°  F.  and  at  the  last  fire-box  from  1400  to  1500°  F.  The 
temperature  of  the  fuel  gases  entering  the  reverberatory  chamber  is 
approximately  1800°  F.  The  ore  never  attains  this  temperature,  because 
it  is  being  continually  stirred  and  advanced,  so  that  only  the  top  layer 
is  momentarily  exposed  to  the  higher  temperatures. 

The  charge  remains  in  the  roasting  furnaces  from  2.5  to  3.0  hours, 
and  on  the  cooling  hearth  from  1.5  to  2.5  hours.  The  bed  is  from  2.5  to 
3.5  in.  thick.  In  some  of  the  furnaces  the  ore  is  rabbled  every  17  seconds; 
in  others  every  35  seconds. 

One  man  attends  a  furnace;  he  usually  has  a  helper  to  wheel  the  coal 
and  ashes.     One  such  helper  attends  several  furnaces. 

The  cost  of  roasting  ranges  from  45  to  55  cents  per  ton,  estimated 
approximately  as  follows,  per  100  tons: 

Coal,  12  tons,  at  .12.00  per  ton, 

Three  f urnacemen,  8-hour  shifts,  at  $2 .  50, 

Coal  and  ash  trammer, 

General  repairs,  oil,  water,  etc., 


Interest  on  investment,  at  6  per  cent.. 

Cost  per  100  tons. 
Cost  per  ton, 

To  this  must  be  added  the  proportional  share  of  administration,  etc., 
which  will  bring  the  cost  per  ton  about  50  cents.  These  estimates  in- 
clude cooling  and  conveying. 

The  furnaces  do  not  give  much  trouble.  It  is  not  unusual  to  have  a 
100-ton  furnace  in  continuous  operation  for  months  without  a  serious 
shut-down  and  no  appreciable  repairs;  with  occasional  changing  of  the 
rabbles  the  roasting  proceeds  indefinitely. 

The  accumulation  of  the  dust  in  the  dust  chambers  is  usually  from 
1.5  to  2.5  per  cent.;  in  some  mills  it  is  re-treated  with  the  ore,  while  in 
others  it  is  shipped  to  the  smelters,  after  briquetting.  The  dust  is  not 
well  roasted,  and  contains  much  undecomposed  pyrites,  and  is  high  in 
soluble  sulphates.  When  the  dust  is  treated  in  the  mills,  it  is  automatic- 
ally again  fed,  with  the  ore,  into  the  furnaces,  and  in  this  way  worked 
up  with  regular  charges. 

The  average  value  of  the  dust  is  higher  than  that  of  the  ore;  this  is 
doubtless  due  to  the  fact  that  in  crushing,  the  sulphides  and  tellurides 


















are  pulverized  more  than  the  quartz.  There  has  not  been  any  appreci- 
able loss  found  by  volatilization  in  oxidizing  roasting. 

Roasting  Arsenical  Sulphide  Ore  at  the  Golden  Gate  Mill,  Mercur, 
Utah. — The  ore  contains  about  5  per  cent,  sulphur,  and  from  4  to  6  per 
cent,  arsenic,  and  small  quantities  of  lead  and  copper.  It  is  crushed  to 
10  mesh,  when  it  is  delivered  to  the  roasters.  The  gold  in  the  ore  is 
found  in  minute  cleavage  planes  and  crevices,  and  is  easily  attacked  by 
the  cyanide  solution;  for  this  reason,  crushing  to  10  mesh  is  required 
only  for  the  sake  of  quick  and  complete  roasting,  and  not  for  the  purpose 
of  facilitating  the  leaching. 

The  plant  consists  of  four  coal-fired  roasters,  125  ft.  long  by  12  1/2  ft. 
wide.  Each  furnace  consists  of  a  single  roasting  hearth,  whose  roof 
serves  as  a  cooling  hearth.  The  ore  is  fed  into  the  lower  level,  and  is 
moved  continuously  in  one  direction  by  travelling  rabbles.  Escaping 
from  the  end,  it  is  elevated  to  the  upper  hearth,  where  it  again  travels  the 
length  of  the  furnace,  exposed  to  the  air. 

The  ore  occupies  about  four  hours  in  roasting,  and  the  same  length 
of  time  in  cooling.  Its  volatile  components  are  reduced  from  5  per  cent, 
sulphur,  and  4  to  6  per  cent,  arsenic,  in  the  raw  state,  to  0.6  per  cent,  sul- 
phur and  0.8  per  cent,  arsenic  in  the  roasted  ore. 

During  the  year  ending  June  30,  1906,  four  roasters  were  in  operation 
for  46  days,  five  roasters  for  237  days,  and  six  roasters  for  82  days. 
The  total  number  of  roaster  days  was,  therefore,  1861.  As  126,358  tons 
of  sulphide  ore  were  roasted,  the  average  work  was  68  tons  per  roaster 
per  day.  The  operation  costs,  including  maintenance  and  repairs,  for 
the  year  1906  were  as  follows: 



Power  and  other  items, 

$126,039  $0,998 

Slack  coal  is  burned  which  costs  15.25  delivered.  The  labor  consists 
largely  of  firemen  who  receive  $2.75  for  eight  hours.  ^ 

Roasting  of  Casilas  Concentrates,  Victoria,  Australia.^ — The  following 
is  an  average  analysis  of  the  concentrate:  Lead,  4.48  per  cent.,;  zinc, 
5.26  per  cent.;  iron,  31.65  per  cent.;  arsenic,  15.16  per  cent.;  sulphur 
31.63  per  cent.;  unestimated,  11.91  per  cent.;  the  copper  rarely  exceeded 
1  or  2  per  cent.  It  was  usual  to  make  from  15  to  20  per  cent,  concen- 
trate, running  from  3  to  4  oz.  of  gold,  and  containing  over  50  per  cent, 
of  the  total  gold  value  of  the  ore  crushed.  The  tailings  from  the  stamp 
battery  and  concentration  tables  are  cyanided,  and  the  concentrate 

^E.  and  M.  J.,  Nov.  10,  1906. 

^Francis  B.  Stephens,  E.  and  M.  J.,  .July  29,  1905. 


Per  ton 








It  wiis  found  that  short  hand-rabbled  reverberatories  were  unsuited 
for  roasting  the  concentrate  on  account  of  the  charge  fusing  too  easily. 
As  the  galena  could  not  all  be  separated,  some  form  of  mechanical 
furnace  was  necessary,  and  two  Edwards'  mechanical  furnaces  with 
60-ft.  hearths  were  installed.  One  man  per  shift  of  8  hours,  attended 
to  all  the  work  of  the  two  furnaces,  with  a  weekly  capacity  of  30  to  35  tons 
each.  The  furnace  man  charged  the  hoppers,  stoked  the  two  furnaces, 
and  looked  after  the  engine  and  dynamo  for  lighting  the  works.  The 
fume  was  lead  into  a  brick  flue  300  ft.  long  and  5  ft.  by  4.  ft.  inside,  with 
a  40-ft.  iron  stack  2  1/2  ft.  in  diameter. 

The  height  of  the  top  of  the  stack  above  the  hearth  of  the  furnaces 
was  70  ft.;  small  dust  chambers  were  built  between  the  ends  of  the  fur- 
nace and  the  flue.  The  iron  stack  did  not  suffer  at  all,  and  acted  as  an 
excellent  arsenic  condenser.  It  was  necessary  to  clean  the  whole  length 
of  the  flue  every  three  months;  about'30  tons  of  deposit  being  obtained. 

The  flue  dust  for  the  first  100  ft.  consisted  partially  of  roasted  con- 
centrate and  arsenic  soot,  assaying  about  3  oz.  in  gold  per  ton,  or  about 
the  same  value  as  the  concentrate  roasted.  The  last  200  ft.  of  the  flue 
contained  arsenic  soot  comparatively  free  from  concentrate,  and  assayed 
7  1/2  dwt.  The  arsenic  at  the  base  of  the  stack  had  to  be  cleared  out 

Just  before  the  finish  of  the  roast,  1  to  1  1/2  per  cent,  salt  was  added 
to  the  concentrate  in  order  to  obtain  a  sweet  roast.  No  evidence  of  loss 
of  gold  by  volatilization  could  be  obtained. 

The  roasted  ore  discharged  into  a  push  conveyor  which  carried  the 
ore  to  a  steel  bucket  elevator;  this  took  it  to  a  cooling  bin  over  the  treat- 
ment vats.  Dry  wood  was  used  as  fuel;  any  green  sticks  getting  in 
generally  had  the  effect  of  throwing  back  the  charge  to  magnetic  oxide. 
Badly  roasted  ore  set  hard  in  the  vats,  while  roasted  ore  did  not.  The 
ore  in  the  hand-rabbled  furnaces  always  roasted  black;  while  in  the 
mechanical  furnaces,  it  roasted  chocolate  color,  but  never  bright  red. 
The  brighter  the  color  obtained  in  roasting,  the  better  the  roast;  although 
the  magnet  failed  to  show  any  difference.  Practically  no  zinc  was  sent 
to  the  flue,  the  heat  not  being  high  enough;  it  was  in  the  roasted  ore 
mostly  as  sulphate  or  basic  sulphate. 

During  the  chlorination  treatment,  the  zinc  was  almost  all  leached 
out  by  the  sulphuric  acid,  but  the  amount  of  zinc  seemed  to  have  no 
influence  on  the  extraction,  other  than  to  prolong  the  period,  through 
packing  of  the  charge  in  the  vats  as  the  zinc  leached  out.  The  chlorine 
solution  used  had  a  strength  of  0.09  to  1.2  per  cent,  of  chlorine,  and  from 
0.5  to  1  per  cent,  sulphuric  acid,  over  and  above  the  amount  required  to 
combine  with  the  beaching  powder.     The  copper  gave  very  little  trouble. 

An  extraction'  of  85  per  cent,  was  obtained  on  well-roasted  ore,  the 
loss  amounting  to  1  to  2  dwt.  per  ton  crushed.     Numerous  experiments 



were  carried  out  to  try  and  better  the  extraction,  but  with  no  success, 
although  they  led  to  a  steady  decrease  of  the  cost  of  chemicals.  Fine 
grinding  after  roasting  gave  no  better  results.  The  cost  of  roasting  the 
concentrate  at  the  rate  of  54  tons  per  week,  was  as  follows : 

General  charges, 

Cost  per  ton  roasted, 



Notwithstanding  the  extremely  refractory  nature  of  the  concentrate, 
over  50  per  cent,  was  saved  by  roasting  and  chlorinating,  over  the  cost 
of  shipping  to  the  smelters.  The  concentrate  sometimes  carried  as  high 
as  20  per  cent.  zinc. 

Roasting  at  Kalgoorlie.' — The  ore  is  crushed  dry  in  Krupp  ball  or  Griffin 
mills  and  roasted  in  Edwards  or  Merton  furnaces.  At  the  Kalgurli  mine 
the  nine  No.  5  Krupp  mills,  eight  of  which  are  in  continuous  use,  are  capable 
of  putting  through  between  1 0,000  and  1 1 ,000  tons  per  month.  These  mills 
maintain  a  duty  of  45  tons  per  24  hours  when  crushing  through  a  37- 
mesh  screen.  The  load  of  balls  weighs  about  2300  lb. ;  one  18-lb.  ball  is 
added  every  day  to  compensate  for  wear  and  tear.  The  mills  run  at 
25  r.  p.  m.  and  require,  including  their  share  of  counter-shaft  friction 
and  dust  fans,  25  h.  p.  The  cost  for  crushing  through  37-mesh  screen 
at  the  South  Kalgurli  is  79  cents  per  ton;  the  cost  at  the  Associated 
and  Associated  Northern,  crushing  through  27-mesh  screen  in  ball  mills, 
is  59  cents. 

The  following  table  gives  the  essential  facts  for  roasting  at  the  Kal- 
goorlie  mines: 



Type  of 

Associated  Northern 



Perse  verence 

Great  Boulder 

Great  Boulder 

South  KalgurU 


Merton.. . 
Merton.. . 
Merton.. . 
Merton.. . 

Area  of 

630  sq.  ft. 
70X9  ft.  6  in. 
63X9  ft.  0  in. 
121X13  ft.  6  in. 
64X6  ft.  6  in. 

422  sq.  ft. 

617  sq.  ft. 

445  sq.  ft. 



(24  hours) 



in  ore 

per  cent. 

Fuel  represents  about  50  per  cent,  of  the  total  roasting  costs. 
Gerard  W.  WilHams,  E.  and  M.  J.  Feb.  15,  1908. 


Fuel  con- 
per  cent, 






Recently '  at  the  Associated,  the  smaller  Merton  furnaces  have  been  re- 
placed by  the  larger  sized  Edwards  furnaces.  The  sulphur  in  the  ore 
averages  about  5.5  per  cent.  The  ore  gets  a  dull  red  heat  about  the 
fifth  rabble  from  the  end.  The  furnaces  average  about  95  tons  per  day 
each.  They  are  motor  driven,  and  use  6  amperes  at  550  volts  each. 
(4.4  h.  p.).  The  end  fire-boxes  are  not  used  much,  just  two  or  three 
logs  are  kept  burning  to  warm  the  air  passing  through  the  fire-bars.  The 
middle  fire-boxes  are  fired  heavily,  and  the  sulphur  continues  to  burn 
till  the  fourth  rabble  from  the  discharge,  namely  22  ft.  from  the  end, 
and  then  discharges  quite  cool.  Fuel  consumption  averages  11  per  cent, 
of  the  roasted  ore.  The  flue  temperature  is  700°  F.  One  man  attends 
two  furnaces.  The  roasting  cost  is  60  cents  per  ton.  About  81,000 
tons  of  ore  and  concentrate  are  roasted  at  Kalgoorlie  monthly. 

'  M.  W.  von  Bernewitz,  Mining  and  Scientific  Press,  May  13,  1911. 




Copper. — Atomic  weight,  63.6;  specific  gravity,  8.94.  Weight  per 
cubic  foot:  cast  copper,  542  lb.;  rolled  copper,  555  lb.  Weight  per 
cu.  in.,  0.32  lb.  Copper  occurs  chemically  as  cuprous  compounds, 
formula  CuA',  or  cupric  compounds,  formula  CuA",  where  AMs  a 
univalent  or  monad  acid  radical,  and  A"  a  bivalent  or  dyad  acid  radical. 
As  a  monad  atom,  copper  has  a  chemical  equivalent  of  63.6,  as  a  dyad 
element  31.8.  The  amounts  of  copper  dissolved  into  or  deposited  from 
a  cupric  or  cuprous  salt  are  proportional  to  the  chemical  equivalent  of 
copper  in  these  two  states  and  to  the  amperes  flowing.  Assuming 
that  one  ampere  liberates  electrolytically  0.00001036  grm.  of  hydrogen 
per  second,  the  amount  of  copper  deposited  by  the  passage  of  one  ampere 
will  be  as  follows: 

Cuprous  compounds 

Cupric  compounds 

One  ampere  per  second  . 
One  ampere  per  m  nute . 
One  ampere  per  hour .  .  . 
One  ampere  per  day .... 
One  ampere  per  year. . . 

0.0006589  grm. 

0.03953  grm. 

2.372  grm. 
56 .  93  grm. 
20 .  78  kilogrm. 

0.0003295  grm. 

0.01977  grm. 

1 .  186  grm. 
28 .  46  grm. 
10 .  39  kilogrm. 

The  melting-point  of  copper  is  1080°  C.  It  is  a  red  metal,  but  thin 
sheets  transmit  a  greenish-blue  light,  and  it  also  shows  the  same  greenish- 
blue  tint  when  in  a  molten  condition.  Of  the  metals  in  ordinary  use, 
only  gold  and  silver  exceed  it  in  malleability.  In  ductility  it  is  inferior 
to  iron  and  cannot  be  so  readily  drawn  into  exceedingly  fine  wire.  Al- 
though it  ranks  next  to  iron  in  tenacity,  its  wire  bears  only  about  half 
the  weight  which  an  iron  wire  of  the  same  size  would  support.  As  a 
conductor  of  heat  it  is  surpassed  only  by  gold.  Next  to  silver  it  is  the 
best  conductor  of  electricity. 

Dry  air  has  no  action  upon  it;  in  moist  air  it  becomes  coated  with  a 
film  of  oxide  which  protects  it  from  further  action  of  air  or  of  water.  It 
forms  a  number  of  very  important  alloys  with  other  metals;  with  tin 
it  forms  bronze;  with  i;inc  and  sometimes  with  small  amounts  of  lead  and 
tin,  it  forms  brass;  and  with  nickel  and  zinc  it  forms  German  silver. 



Copper  which  has  become  hardened  by  mechanical  work  may  be 
again  made  malleable  by  heating.  The  boiling-point  of  copper  is  about 
2000°  C.  Molten  copper  has  a  great  tendency  to  dissolve  hydrogen, 
carbonic  oxide,  and  sulphur  dioxide,  which  it  evolves  again  on  solidi- 
fying. Aluminum,  cobalt,  nickel,  zinc,  cadmium,  tungsten,  molyb- 
denum and  iron,  are  more  or  less  readily  dissolved  by  it,  as  also  are  cup- 
rous oxide,  sulphide,  and  phosphide,  and  the  arsenides,  arsenates,  anti- 
monides  and  antimonates. 

On  heating  copper  to  a  low  red  heat,  far  below  its  melting  point,  it 
becomes  covered  with  a  film  or  scale  which  consists  of  a  mixture  of  the 
cuprous  and  cupric  oxides. 

Copper  exhibits  a  greater  affinity  for  sulphur  than  do  any  of  the  other 
metals.  It  also  unites  directly  with  the  metalloids,  excepting  hydrogen, 
nitrogen,  and  carbon. 

The  best  solvents  for  copper  are  nitric  acid,  concentrated  sulphuric 
acid,  and  aqua  regia.  Hydrochloric  acid  and  dilute  sulphuric  acid  only 
dissolve  the  metal  when  air  or  some  other  oxidizing  substance  is  present; 
under  these  conditions  it  is  more  readily  soluble  in  dilute  hydrochloric 
acid  than  in  dilute  sulphuric  acid. 

Cupric  chloride  acts  on  metallic  copper  to  produce  cuprous  chloride: 
Cu  +  CuCl2  =  2CuCl. 

Cuprous  oxide  has  the  property  of  mixing  with  molten  copper  in  all 
proportions.  Small  amounts  of  cuprous  oxide  have  no  injurious  effect 
upon  it,  but  large  quantities  make  it  cold-short,  and  when  a  certain 
limit  is  exceeded,  also  red-short.  Copper  containing  about  2  per  cent, 
cuprous  oxide  is  still  as  fit  for  ordinary  use  as  ordinary  cast-refined 

Cathode  copper  is  exceedingly  pure,  usually  about  99.93  per  cent, 
copper,  with  hydrogen  as  the  chief  impurity.  Objectionable  cathode 
impurities  are  of  two  classes — those  which  depress  the  electrical  conduc- 
tivity and  those  which  make  the  metal  brittle.  Arsenic  and  antimony 
represent  the  first  class;  tellurium  and  lead  the  second.  Good  cathode 
copper  should  show  but  a  few  thousandths  of  a  per  cent,  of  arsenic  and 
antimony.  Experiments  have  indicated  that  it  takes  but  0.0013  per 
cent,  of  arsenic  or  0.007JL  per  cent,  of  antimony  to  lower  the  conduc- 
tivity 1  per  cent.  Any  conductivity  troubles  in  electroljrtic  copper 
can  almost  invariably  be  traced  to  the  presence  of  undue  amounts  of  one 
or  both  of  these  elements.  Impurities  of  the  brittle-making  class  are 
rarely  met  with,  and  if  present  are  due  to  mechanical  contamination 
of  the  cathode,  either  in  the  bath  or  in  the  subsequent  furnace  treatment. 

Influence  of  Impurities  on  the  Properties  of  Copper. 

Arsenic  and  Antimony. — Hampe,  in  1892,  found  that  0.5  per  cent, 
arsenic  produces  no  bad  results  and  that  even  when  the  percentage  was 


increased  to  1  per  cent,  only  a  slight  degree  of  red-shortness,  but  no  cold- 
shortness  could  be  noticed.  He  found  that  copper  with  0.8  per  cent, 
arsenic  could  be  drawn  into  the  finest  wire.  Stahl,  in  1886,  stated  that 
a  small  percentage  of  arsenic  prevents  copper  from  becoming  porous. 
Hiorns,  in  1906,  showed  that  copper  with  arsenic  up  to  0.4  per  cent,  was 
very  malleable  when  cold;  that  with  0.2  per  cent,  each  of  arsenic  and  anti- 
mony, the  same  is  true;  and  that  arsenic  in  the  presence  of  antimony 
makes  the  copper  more  malleable  than  it  is  with  antimony  alone,  though 
antimony  when  not  above  0.2  per  cent,  only  slightly  impairs  the  mal- 
leability of  copper.  He  adds  that  arsenic  in  copper  is  highly  beneficial 
because  it  deoxidizes  cuprous  oxide,  which  tends  to  destroy  the  mallea- 
bility of  copper.     Johnson,  in  1906,  stated  that  cast  copper  with  0.5  per 




6  40 











ire  Cop 



\     Te 


are  13.4 




- — 


1  2 

Percentage  of  Arsenic 

Pjq  43 Conductivity  diagram  of  pure  electrolytic  copper  with  arsenic  and  antimony 

as  alloys. 

cent,  arsenic  has  a  tensile  strength  of  10  long  tons  per  square  inch,  and 
a  24  per  cent,  elongation.  After  forging,  the  tensile  strength  was  raised 
to  12.75  tons,  and  the  elongation  to  35  per  cent.  Upon  rolling,  the 
tensile  strength  became  14  tons  and  the  elongation  48  per  cent.;  and 
finally  upon  being  highly  wrought  and  cold  drawn,  the  tensile  strength 
of  the  same  cast  copper  was  raised  to  15.9  tons,  and  the  elongation  varied 


from  24  to  50  per  cent,  while  the  specific  gravity  was  increased  from 
8.83  for  copper  in  the  cast  state  to  8.866  when  the  metal  was  wrought. 

H.  S.  Hiorns  and  S.  Lamb'  prepared  alloys  consisting  of  pure  elec- 
trolytic copper  and  arsenic  and  antimony  in  quantities  va,rying  from 
0.05  to  3.5  per  cent.  These  alloys  were  drawn  into  wires  0.0325  in.  in 
diameter,  and  were  tested  for  conductivity,  shown  by  Fig.  43. 

A.  H.  Hiorns  found  that  with  between  0.5  and  1  per  cent,  of  arsenic 
the  malleability  seemed  to  diminish,  but  with  over  1  per  cent,  and  up  to" 
2  and  3  per  cent,  arsenic,  the  copper  rolled  perfectly  and  was  harder  than 
pure  copper.  With  less  than  0.5  per  cent.,  arsenic  the  copper  should 
be  less  malleable  when  cooled  slowly  than  when  cooled  quickly.  With 
a  certain  small  quantity  of  arsenic  introduced  into  copper  the  first  small 
portions  appear  to  act  by  reducing  the  cuprous  oxide,  the  remainder 
retaining  the  metallic  form  and  toughening  the  copper. 

Arsenic  invariably  improves  the  forging  properties  where  added  to 
impure  copper.  Arsenical  copper  is  largely  specified  for  in  materials, 
such  as  locomotive  and  boiler  tubes,  which  are  required  to  withstand  high 
temperatures,  since  mechanically  hardened  arsenical  copper  is  not 
softened  at  so  low  an  annealing  temperature  as  pure  electrolytic  copper, 
which  has  undergone  the  same  treatment. 

Arsenic  appears  to  improve  the  hot-working  properties  of  copper 
vitiated  by  traces  of  bismuth. 

Bismuth. — Bismuth  is  the  most  injurious  impurity  in  copper,  as 
very  small  quantities  render  the  copper  unworkable.  According  to 
Hampe,  copper  containing  as  little  as  0.02  per  cent,  of  bismuth  is  red- 
short,  and  0.05  per  cent.,  cold-short.  With  0.1  per  cent,  the  copper 
crumbles  under  the  hammer  at  a  red  heat.  The  presence  of  certain 
proportions  of  arsenic  and  antimony  somewhat  counteracts  the  tend- 
ency of  bismuth  to  produce  cold-shortness. 

Lead.  — ^Lead  can  be  melted  with  copper  in  all  proportions,  but  the 
greater  part  of  it  can  be  liquated  out  of  the  alloy  by  a  gentle  heat.  Lead 
is  to  be  found  in  all  ordinary  commercial  copper,  but  it  is  not  desirable 
in  any  proportion  over  0.10  per  cent.,  and  the  lower  the  proportion 
under  0.10  per  cent,  the  better.  According  to  Hampe's  experiments, 
0.15  per  cent,  of  lead  does  not  affect  the  malleability  of  copper  in  any 
way;  with  0.3  per  cent,  of  lead  it  becomes  slightly  red-short,  and  with 
0.4  per  cent,  slightly  cold-short.  With  1  per  cent,  it  is  unworkable. 
The  lead  reduces  the  strength,  ductility  and  toughness  of  copper. 
The  solvent  action  of  copper  for  lead  is  very  small.  The  addition  of 
lead  to  copper  has  the  effect  of  lowering  the  affinity  of  copper  for 
reducing  gases. 

Iron. — Iron  forms  no  true  alloy  with  copper;  small  admixtures  of 
iron  such  as  are  contained  in  many  varieties  of  copper  have  no  injurious 
•Min.  Ind.,  Vol.  XVIII,  1909. 


effects  upon  it.  It  occurs  in  refined  copper  in  the  merest  traces,,  which  are 
quite  harmless.  When  all  the  sulphur  has  gone  from  a  charge  of  copper, 
in  refining,  a  sample  taken  from  the  furnace  shows  an  unblistered  surface, 
and  is  said  to  be  "set-copper."  At  this  point  all  the  iron  has  been  elimi- 
nated except  the  merest  trace. 

Iron  acts  as  a  deoxidizer  when  added  to  copper.  Copper  containing 
only  1  per  cent,  iron  is  rendered  feebly  magnetic,  will  forge  well  at  a  red 
heat,  is  quite  malleable,  tough  and  strong,  even  in  the  presence  of  arsenic. 
It  lacks  the  fluidity  of  pure  copper  when  poured  at  the  same  temperature. 

Nickel. — Nickel  alloys  with  copper  in  all  proportions.  Traces 
of  nickel  in  copper  are  beneficial,  imparting  strength  and  toughness. 
Nickel  must  be  kept  low  if  arsenic  is  present,  0.10  per  cent,  being  quite 
sufficient  to  harden  arsenical  copper  which  has  to  withstand  severe 
working.  Below  0.05  per  cent.,  even  in  the  presence  of  arsenic,  its 
effect  upon  the  physical  properties,  excepting  electrical  conductivity, 
may  be  considered  insignificant. 

Cobalt.— Little  is  known  about  the  influence  of  this  metal  on  copper. 
According  to  F.  Johnson,  it  toughens  and  strengthens  copper  when 
present  up  to  at  least  1  per  cent,  without  impairing  its  hot-working 
qualities.  Probably  it  acts  very  similar  to  nickel,  conferring  greater 
durability  at  high  temperatures,  while  toughening,  hardening,  and 
strengthening  the  copper  in  the  cold.  Cobalt,  in  the  low  percentages  in 
which  it  is  found  in  copper,  is,  if  anything,  beneficial,  and,  moreover,  it 
does  not  disagree  with  arsenic  to  so  great  an  extent  as  nickel  does. 

Tin. — ^Tin  hardens  copper,  more  than  any  other  element.  It  occurs 
very  rarely  in  commercial  copper  being  readily  eliminated  during  the 
process  of  reducing  the  copper.  Low  percentages  of  tin  improve  the 
tensile  strength,  ductility  and  resistance  to  corrosion,  and  maintain 
these  improvements  at  high  temperatures,  but  the  natural  softness  of 
copper  and  its  red  color  are  both  materially  removed.  Its  malleability 
is  also  decreased. 

Tellurium. — ^A  few  thousandths  of  1  per  cent,  of  tellurium  renders 
copper  appreciably  red-short;  but  very  little  is  known  of  the  effect  of 
tellurium  in  commercial  copper. 

Sulphur  rarely  occurs  in  more  than  harmless  traces  in  commercial 
copper,  yet  it  may  occur  to  the  extent  of  0.03  per  cent.  (SOj)  in  electro- 
lytic copper  having  a  conductivity  of  102.2  per  cent.  It  derives  its 
origin  from  the  incomplete  removal  of  sulphur  from  the  sulphate  liquor 
in  which  the  copper  cathodes  were  deposited,  and  from  the  sulphurous 
gases  from  the  fuel  of  the  reverberatory  refining  furnace  where  it  is 
partially  dissolved  by  the  molten  metal,  as  they  pass  over  it  on  their  way 
to  the  flue.  In  the  first  case  the  sulphate  would  probably  be  reduced  to 
sulphide  by  the  reducing  action  of  poling,  and  in  the  second  case,  sulphur 
dioxide  would  be  absorbed  and  retained  as  such.     Cuprous  sulphide  is  a 


highly  undesirable  constituent  of  copper,  and  its  presence  in  any  alloy 
would  be  detrimental.  Hampe  finds  that  copper  with  0.25  per  cent, 
sulphur  is  still  moderately  malleable,  but  with  0.5  per  cent,  it  becomes 
very  cold-short,  although  not  red-short. 

Carbon  is  not  at  all  absorbed  by  copper. 

Pure  copper  or  copper  of  more  than  usual  purity,  assaying,  say,  99.8 
per  cent.,  is  inferior  to  impure  copper  in  mechanical  properties,  dura- 
bility, and  resistance  to  corrosion. 

Copper,  when  cast  in  moulds,  has  the  property  of  rising  and  becoming 
porous.  Sound  castings  can  only  be  obtained  by  means  of  special 
precautions,  such  as  pouring  at  the  lowest  possible  temperature,  or 
pouring  in  an  atmosphere  of  carbon  dioxide. 

Cupric  Carbonate. — The  normal  carbonate  has  not  been  obtained. 
The  two  most  important  basic  carbonates  are: 

(1)  CuC03,Cu(OH)2,  which  occurs  native  as  malachite. 

(2)  2CuC03,Cu(OH)j,  which  occurs  native  as  azurite. 

The  first  is  obtained  when  sodium  carbonate  is  added  to  a  solution 
of  copper  sulphate.  When  these  carbonates  are  slowly  heated  to  220°  C. 
the  carbonate  is  slowly  converted  into  black  cupric  oxide. 

The  carbonates  of  copper  are  readily  soluble  in  dilute  sulphuric,  hydro- 
chloric, sulphurous  and  nitric  acids.  They  are  also  readily  soluble  in 
ammonia  and  ammonia  salts.  They  are  partially  soluble  in  sodium 
carbonate  and  in  solutions  of  potassium  cyanide. 

-Cupric  Nitrate,  Cu(N03),3H20. — Cupric  nitrate  may  be  obtained  by 
the  action  of  nitric  acid  upon  cupric  oxide,  hydroxide,  carbonate,  or 
the  metal  itself.  Copper  is  soluble  in  nitric  acid,  in  all  of  its  mineralogical 
combinations.     Sulphides  are  decomposed,  as  solution  takes  place. 

Cupric  nitrate  is  very  easily  soluble  in  water. 

Cupric  Oxide,  CuO  (Black  Oxide  of  Copper). — Cupric  oxide  occurs  in 
nature  as  the  rather  rare  mineral,  tenorite.  It  may  be  prepared  artifi- 
cially by  continued  ignition  of  copper  in  contact  with  air;  by  exposing 
cupric  sulphate  to  an  intense  red  heat,  or  the  carbonate,  nitrate,  or 
hydroxide  to  a  moderate  heat. 

When  caustic  potash  or  soda  is  added  by  drops  to  a  boiling  solution 
of  cupric  salts  till  the  acid  is  saturated  the  whole  of  the  copper  is  precipi- 
tated as  anhydrous  black  oxide,  which  may  be  freed  from  potash  or 
soda  by  boiling  with  water. 

Cupric  oxide  is  a  black  powder,  which  rapidly  absorbs  moisture 
from  the  air.  When  heated  it  first  cakes  together  and  finally  fuses, 
giving  up  part  of  its  oxygen,  and  leaving  a  residue  consisting  of  CuO,2Cu20. 
When  heated  with  charcoal,  or  in  a  stream  of  carbon  monoxide,  marsh 
gas  or  hydrogen,  it  is  reduced  to  the  metallic  state. 


When  the  cupric  oxide  is  gently  heated  with  metallic  copper,  it  is 
converted  into  cuprous  oxide. 

A  mixture  of  cupric  oxide  with  excess  of  sulphur  is  resolved  at  a  red 
heat  into  cuprous  sulphide,  sulphur  dioxide  and  a  trace  of  cupric  sulphate. 
If  on  the  contrary,  the  cupric  oxide  is  in  excess,  cuprous  oxide  and  cupric 
sulphate  are  produced,  and  with  only  a  trace  of  sulphur  dioxide,  except- 
ing that  when  the  heat  is  raised  to  the  point  at  which  the  cupric  sulphate 
is  decomposed. 

Cupric  oxide  has  a  strong  affinity  for  acids,  dissolving  in  them  easily. 
It  is  soluble  in  sulphurous  acid.  It  is  insoluble  in  ammonia,  but  dissolves 
on  the  addition  of  a  few  drops  of  acid  or  ammonium  carbonate.  It  is 
insoluble  in  dilute,  but  soluble  in  warm  concentrated  caustic  soda  or 
potash.  Ferrous  chloride  converts  cupric  oxide  into  cuprous  and  cupric 
chlorides,  with  the  formation  of  ferric  oxide.  Ferric  chloride  converts 
cupric  oxide  into  cupric  chloride,  with  the  formation  of  ferric  oxide. 

Cupric  oxide  is  reduced  to  cuprous  oxide  at  1050°  C. 

Cuprous  Oxide,  CU2O  (Red  Oxide  of  Copper). — Cuprous  oxide  occurs 
native  as  cuprite,  the  red  pxide  of  copper.  It  is  formed  when  finely 
divided  copper  is  gently  heated  in  a  current  of  air  or  when  a  mixture  of 
cuprous  chloride  and  sodium  carbonate  is  gently  heated  in  a  covered 

Cuprous  oxide  is  reduced  to  the  metallic  stage  by  gentle  ignition  with 
charcoal  or  hydrogen. 

Cuprous  oxide  is  insoluble  in  water ;  it  is  converted  into  cuprous  chloride 
by  hydrochloric  acid.  Nitric  acid  converts,  it  into  cupric  nitrate  with 
the  evolution  of  oxide  of  nitrogen.  When  acted  upon  by  dilute  sulphuric 
acid,  it  is  partly  reduced  to  metallic  copper  and  partly  oxidized  into 
copper  sulphate.  When  heated  with  strong  acid  it  is  entirely  oxidized 
to  sulphate. 

When  copper  is  oxidized  with  a  considerable  quantity  of  oxygen  at 
a  high  temperature,  it  forms  cupric  oxide  (CuO).  If  the  ignition  be 
carried  further,  cuprous  oxide,  CujO,  may  be  formed  from  the  CuO. 
The  cuprous  oxide  is  not  as  readily  soluble  as  the  cupric  oxide,  and  it 
may  be  partly  for  this  reason  that  copper  sulphides  roasted  at  a  high 
temperature  do  not  give  a  good  extraction  of  the  contained  copper. 

Cuprous  oxide  fuses  at  a  red  heat. 

When  heated  with  acids  coprous  oxide  forms  a  solution  of  a  cupric 
salt  and  metallic  copper;  for  example, 

Cu20  +  H2SO,  =  CuSO,  +  Cu  +  H20. 

However,  strong  hydrochloric  acid  does  not  deposit  metallic  copper 
on  dissolving  cuprous  oxide,  which  is  due  to  the  fact  that  the  cuprous 
chloride  formed  is  soluble  in  strong  hydrochloric  acid. 

Cupric  Sulphate,  CuSO^jSH^O.— Copper  sulphate  may  be  formed 



by  applying  dilute  sulphuric  acid  to  copper  oxide,  when  the  sulphate 
crystallizes  out  on  cooling;  by  heating  metallic  copper  with  concentrated 
sulphuric  acid,  whereupon  sulphur  dioxide  is  evolved,  and  anhydrous 
cupric  sulphate  is  precipitated  as  a  white  powder,  mixed  with  a  brown 
mass  of  cuprous  and  cupric  sulphides;  on  digesting  this  mass  with  hot 
water,  the  cupric  sulphate  dissolves,  and  may  be  crystallized  out  of  the 

On  roasting,  the  sulphide  ores  of  copper  are  converted  into  cupric 
oxide  and  cupric  sulphate.  When  water  is  applied  to  the  roasted  ore, 
the  copper  sulphate  is  dissolved;  by  evaporation  of  the  water,  the  copper 
sulphate  crystallizes  out  of  the  solution. 

At  100°  C.  copper  sulphate  loses  4  molecules  of  water,  and  at  200°  it 
loses  all  its  water.  At  a  bright  red  heat  it  decomposes  into  copper 
oxide  and  sulphuric  acid.  When  heated  with  carbon  at  a  dark  red  heat 
the  copper  is  separated,  with  the  formation  of  carbonic  acid  and  sul- 
phur dioxide. 

From  solutions  of  copper  sulphate,  the  copper  is  precipitated  by  means 
of  iron,  aluminum,  and  zinc,  as  metallic  copper;  with  hydrogen  sulphide  or 
the  sulphide  of  the  alkali  metals,  it  is  precipitated  as  the  cupric  sulphide 
(CuS) .  By  electrolysis,  copper  is  deposited  from  copper  sulphate  solutions 
at  the  cathode  and  acid  liberated  at  the  anode.  If  at  the  anode,  ferrous 
sulphate  is  present  in  the  solution,  it  is  converted  into  ferric  sulphate. 

The  crystallized  copper  sulphate  dissolves  in  3  1/2  parts  of  cold  water, 
and  in  much  smaller  quantities  of  boiling  water. 

In  100  parts  of  water,  at  the  following  temperatures 


Parts  CuSOi 




































Per  cent,  copper 












All  of  the  chlorides  have  the  faculty  of  converting  copper  sulphate,  in 
solution,  into  the  chloride.     Hydrochloric  acid  dissolves  copper  sulphate 


with  considerable  reduction  of  temperature,  forming  a  green  liquid, 
which  when  evaporated  forms  crystals  of  cupric  chloride. 

When  excess  of  ammonia  is  added  to  a  solution  of  copper  sulphate, 
a  deep  blue  solution  is  formed  having  the  composition  CuS04,H20,4NH3. 

Cupric  and  ferrous  sulphates  cannot  be  entirely  separated  by  crystal- 
lization, as  a  solution  of  these  salts  deposits  a  double  sulphate  of  the  two 
metals.  If,  however,  the  amount  of  iron  present  is  comparatively  small, 
the  first  crop  of  crystals  obtained  is  moderately  pure  copper  sulphate. 

Cupric  Chloride,  CuClj. — Cupric  chloride  may  be  obtained  in  the 
anhydrous  condition  by  the  combustion  of  copper  in  an  atmosphere  of 
chlorine  gas;  copper  filings  or  copper  foil  introduced  into  dry  chlorine 
takes  fire  spontaneously,  and  burns  with  a  greenish  light,  producing 
a  mixture  of  cupric  and  cuprous  chlorides,  and  if  the  chlorine  is  in  excess 
the  cuprous  chloride  is  slowly  converted  into  the  cupric  chloride.  It 
is  also  produced  when  compounds  of  copper  are  roasted  with  salt  or 
other  chlorides. 

In  the  wet  way  it  is  formed  when  copper  is  dissolved  in  nitro-hydro- 
chloric  acid  (aqua  regia),  or  when  cupric  oxide,  carbonate,  or  hydroxide 
are  dissolved  in  hydrochloric  acid.  Cupric  chloride  is  readily  soluble 
in  water,  forming  a  deep  green  solution,  which  on  being  largely  diluted, 
turns  blue.  The  salt  crystallizes  in  green  rhombic  prisms,  with  2H2O, 
giving  the  composition  of  the  crystals  as  CuCl2,2H20.  When  heated 
to  200°  C.  it  loses  its  water  of  crystallization,  and  at  a  dull  red  heat  is 
converted  into  cuprous  chloride,  with  evolution  of  chlorine. 

With  copper  oxides  cupric  chloride  combines  in  various  proportions 
to  form  oxychlorides.  From  solutions  of  cupric  chloride  metallic  copper 
is  precipitated  by  iron,  aluminum,  and  zinc.  With  hydrogen  sulphide 
and  the  sulphides  of  the  alkali  metals  and  earths  the  copper  is  precipi- 
tated as  cupric  sulphide,  (CuS).  Calcium  hydroxide,  or  lime,  precipi- 
tates copper  as  the  hydroxide,  which  on  heating,  is  converted  into  the 
oxide.  By  passing  sulphur  dioxide  into  a  solution  of  cupric  chloride,  the 
copper  is  precipitated  as  the  cuprous  chloride. 

By  electrolysis,  copper  is  deposited  from  cupric  chloride  solutions 
at  the  cathode,  while  chlorine  is  liberated  at  the  anode. 

In  100  parts  of  water,  at  the  following  temperatures 


Parts  CuClj  Per  cent,  copper 







70 . 6  !  33 . 6  per  cent. 

76 . 2  I  36 . 3  per  cent. 


100  parts  of  water,  saturated  with  CuClj,  contains,  at  the  following 
temperatures : 






17.0           1             62 



Parts  CuClj  Per  cent,  copper 

41.4  i  19.7  per  cent. 

43.1  20.4  per  cent. 

44.6  21.2  per  cent. 

100  grm.  of  water  dissolve  121.4  grm.  of  CUCI2+2H2O,  at  16.1°  C. 

Cupric  chloride  is  not  decomposed  by  cold  sulphuric  acid. 

It  is  soluble  in  solutions  of  ammonium  chloride,  and  very  soluble  in 
concentrated  solutions  of  common  salt.  It  is  less  soluble  in  concentrated 
solutions  of  hydrochloric  acid  than  in  dilute  solutions. 

With  ammonia  cupric  chloride  forms  a  deep  blue  solution  having  the 
composition  CuCl2,4NH3,H20. 

Cuprous  Chloride,  CujClj. — Cuprous  chloride  may  be  obtained  by 
dissolving  cuprous  oxide  in  hydrochloric  acid.  It  is  more  readily  pre- 
pared by  boiling  a  solution  of  cupric  chloride  in  hydrochloric  acid,  with 
copper  foil  or  copper  turnings.  The  nascent  hydrogen,  liberated  by  the 
action  of  hydrochloric  acid  upon  the  copper,  reduces  the  cupric  chloride  to 
the  cuprous  chloride.  The  liquid  is  then  poured  into  water,  which  causes 
the  precipitation  of  the  cuprous  chloride  as  a  white  crystalline  powder. 
A  mixture-  of  zinc  dust  and  copper  oxide  added  to  strong  hydrochloric 
acid,  also  yields  cuprous  chloride,  the  nascent  hydrogen  in  this  case 
being  derived  from  the  zinc,  and  this  causes  the  reduction  of  cupric 
chloride  formed  by  the  action  of  the  acid  upon  the  cupric  oxide. 

Cuprous  chloride  may  be  formed  by  heating  the  cupric  chloride  to 
a  dull  red  heat. 

Cuprous  chloride  melts  somewhat  below  a  dull  red  heat,  and  when 
slowly  cooled,  solidifies  in  a  translucent  yellow  mass.  In  closed  vessels 
it  does  not  volatilize,  even  when  strongly  heated,  but  if  heated  in  the 
air  it  goes  off  in  white  vapor.  When  exposed  to  the  air  in  a  dry  state 
it  slowly  absorbs  moisture  and  turns  green;  in  the  moist  state  it  is 
quickly  turned  into  a  green  mass,  of  oxychloride  of  copper,  CuClj,- 
3CuO,4H20.     This  compound  occurs  native  as  the  mineral  Atacamite. 

Cupric  chloride,  CuClj,  when  ignited  gives  cuprous  chloride,  and 
therefore  cuprous  chloride  is  always  formed  when  copper  enters  into 
reaction  with  chlorine  at  a  high  temperature.  The  green  solution  of 
cupric  cliloride  is  decolorized  by  metallic  copper,  cuprous  chloride  being 
formed;  but  this  reaction  is  only  accomplished  with  ease  when  the  solu- 


tion  is  very  concentrated  and  in  the  presence  of  an  excess  of  hydrochloric 
acid  to  dissolve  the  cuprous  chloride.  The  addition  of  water  precipitates 
cuprous  chloride. 

Many  reducing  agents  which  are  capable  of  taking  up  half  the  oxygen 
from  cupric  oxide  are  able,  in  the  presence  of  hydrochloric  acid,  to  form 
cuprous  chloride;  sulphur  dioxide,  SOj,  acts  in  this  manner.  The  usual 
method  of  preparing  cuprous  chloride  consists  in  passing  sulphur  dioxide 
into  a  strong  solution  of  cupric  chloride. 

Cuprous  chloride  forms  colorless  cubic  crystals  which  are  insoluble 
in  water.  Under  the  action  of  oxidizing  agents,,  it  passes  into  cupric 
salts  and  it  absorbs  oxygen  from  the  moist  air,  forming  cupric  oxychloride. 

From  solutions  of  cuprous  chloride,  metallic  copper  is  precipitated 
by  iron,  aluminum,  and  zinc.  Hydrogen  sulphide  and  the  sulphides 
of  the  alkali  metals  and  earths,  precipitate  the  copper  as  cupric 
sulphide,  CuS. 

By  electrolysis,  copper  is  deposited  at  the  cathode,  while  chlorine  is 
liberated  at  the  anode.  If  univalent  salts  are  present  in  the  anode 
solution,  these  will  be  converted  into  bivalent  salts  by  the  action  of  the 
liberated  chlorine. 

Milk  of  lime,  added  to  a  hot  solution  of  cuprous  chloride,  precipitates 
the  copper  as  cuprous  oxide. 

Cuprous  chloride  is  insoluble  in  water,  but  dissolves  in  hydrochloric 
acid,  ammonia,  and  alkaline  chlorides. 



Saturated  sodium  chloride  solution  dissolves  at 

Degrees  C. 

Degrees  P. 

Cuprous  chloride,  CU2CI2 

Metallic  Cu 



16.9  per  cent. 

11.9  per  cent. 

8 . 9  per  cent. 

10.76  per  cent. 
7 .  65  per  cent. 
5 .  73  per  cent. 

15  per  cent. 

VaCl-Aq.  dissolves  at 


*     194 

10.3  per  cent. 
6 . 0  per  cent. 
3 . 6  per  cent. 

6.62  per  cent. 
3 .  86  per  cent. 
2 .  31  per  cent. 


5  per  cent.  NaCl-Aq.  dissolves  at 



2 . 6  per  cent. 
1 . 1  per  cent. 

1 .  67  per  cent. 
0 .  71  per  cent. 

Cuprous  chloride,  when  melted,  conducts  the  electric  current  very 
well,  copper  separating  out  as  fine  leaves.  The  melt  cannot  be  heated 
to  the  melting  point  of  copper  and  the  copper  obtained  liquid,  because 
the  cuprous  chloride  vaporizes  too  easily. 

Cupric  Silicate,  CuSi03  +  2H20. — Silicate  of  copper  occurs  native  as 
chrysocoUa,  CuSi03+2H20,  and  dioptase,  CuH2Si04.  Chrysocolla  is 
soluble  in  dilute  hydrochloric  acid,  leaving  a  residue  of  silica.  Dioptase 
is  soluble  in  nitric  and  hydrochloric  acids,  or  ammonia,  with  separation 
of  gelatinous  silica.     It  is  not  attacked  by  caustic  alkalies. 

Cuprous  Sulphide,  CujS. — There  are  two  sulphides  of  copper,  corre- 
sponding to  the  two  oxides;  the  cuprous  sulphide,  CujS,  and  the  cupric 
sulphide,  CuS. 

The  cuprous  sulphide,  when  heated  at  a  comparatively  low  temper- 
ature, loses  one-half  of  its  sulphur  and  is  converted  into  the  cupric 

Cuprous  sulphide  occurs  in  nature  as  copper  glance,  or  chalcocite. 
It  is  produced  artificially  when  copper  burns  in  sulphur  vapor,  or  when 
an  excess  of  copper  filings  is  heated  with  sulphur. 

It  is  not  decomposed  out  of  contact  with  the  air;  but  if  air  has  access 
to  it,  combustion  takes  place,  and  sulphur  trioxide  and  cupric  oxide  are 
produced.  When  heated  to  redness  in  a  current  of  aqueous  vapor,  it  is 
but  slightly  decomposed,  but  at  a  white  heat,  it  yields  large  quantities  of 
hydrogen  and  hydrogen  sulphide  together  with  sublimed  sulphur,  and 
the  copper  is  completely  reduced  to  the  metallic  state.  It  is  not  altered 
by  ignition  in  a  stream  of  hydrogen. 

It  is  not  decomposed  by  chlorine  gas  at  ordinarj^  temperatures;  very 
slowly  when  heated.  It  dissolves  with  difficulty  in  strong  boiling  hydro- 
chloric acid.  In  heated  nitric  acid  it  dissolves  with  separation  of  sulphur, 
whereas  cold  nitric  acid  dissolves  one-half  the  copper,  and  leaves  the 
cupric  sulphide.  Cuprous  sulphide,  ignited  with  cuprous  oxide,  is 
easily  converted  into  sulphur  dioxide  and  copper  or  cuprous  oxide.  It 
is  not  dissolved  by  sulphuric  acid.  It  is  slowly  acted  upon  by  solutions 
of  ferric  chloride  and  of  ferric  sulphate.  Cuprous  sulphide  melts  at 
1127° C. 

Cupric  Sulphide,  CuS. — Cupric  sulphide  is  met  with  in  nature  as  the 
mineral  covelite  (blue  copper).  It  is  obtained  artificially  when  either 
copper  or  cuprous  sulphide  is  heated  with  sulphur  to  a  temperature  not 
beyond  114°  C;  so  obtained,  the  compound  is  blue.     As  a  black  precip- 


itate,  it  is  formed  when  hydrogen  sulphide  is  passed  into  solutions  of 
cupric  salts. 

Treated  with  hot  nitric  acid  the  copper  is  oxidized,  part  of  the  copper 
is  converted  into  sulphate  and  the  rest  separated,  so  that  the  resulting 
solution  contains  both  nitrate  and  sulphate  of  copper.  Hot  concentrated 
hydrochloric  acid  slowly  converts  it  into  cupric  chloride,  with  evolution 
of  hydrogen  sulphide  and  separation  of  sulphur.  Cupric  sulphide  decom- 
poses silver  salts,  the  copper  dissolving  and  the  sulphide  of  silver  being 
precipitated.  It  is  insoluble  in  dilute  sulphuric  acid,  caustic  alkalies,  and 
fixed  alkaline  sulphides.     It  is  slightly  soluble  in  ammonium  sulphide. 

Cupric  Hydroxide,  Cu(0H)2. — Cupric  hydroxide  is  a  pale  blue  pre- 
cipitate produced  when  sodium  or  potassium  hydroxide  is  added  in 
excess  to  a  solution  of  a  copper  salt.  The  compound,  when  washed, 
may  be  dried  at  100°  C,  without  parting  with  water;  but  if  the  liqu 
in  which  it  is  precipitated  be  boiled,  the  compound  blackens,  apd  is  con- 
verted into  a  hydrate  having  the  composition  Cu(0H)2,  2CuO.  Cupric 
hydroxide  dissolves  in  ammonia,  forming  a  deep  blue  liquid.  It  is  very 
soluble  in  acids.  It  is  changed,  by  standing,  to  the  black  compound, 
Cu302(OH)2  and  by  boiling  to  cupric  oxide,  CuO. 

Ammonium  carbonate,  like  ammonium  hydroxide,  precipitates  the 
cupric  hydroxide  and  redissolves  it  to  a  blue  solution.  Carbonates  of 
the  fixed  alkali  metals,  as  potassium  and  sodium  carbonate,  precipitate 
the  greenish-blue  carbonate,  Cu2(OH)2C03,  which  is  converted  by  boiling 
to  the  black,  basic  hydroxide,  and  finally  to  the  black  oxide. 

From  the  blue  ammoniacal  solutions  a  concentrated  solution  of  a 
fixed  alkali  precipitates  the  blue  hydroxide,  changed  on  boiling  to  the 
black  oxide,  CuO. 

Cupric  hydroxide  is  soluble  in  a  solution  of  cane  sugar  in  the  presence 
of  an  alkali  or  alkaline  earth.  It  is  somewhat  soluble  in  the  caustic 
alkalies,  and  very  soluble  in  ammonia. 

Copper  Cyanides. — ^Potassium  cyanide  forms,  with  copper,  the  yel- 
lowish-green cupric  cyanide,  Cu(CN)2,  soluble  in  excess,  with  the  forma- 
tion of  the  double  cyanide,  2KCN,  Cu(CN)2,  unstable,  changing  in  whole 
or  in  part  to  cuprous  cyanide.  The  potassium  cyanide  also  dissolves 
cupric  oxide,  hydroxide,  carbonate,  sulphide,  etc.,  changing  rapidly  to 
cuprous  cyanide  in  solution  in  the  alkali  cyanide. 

Potassium  ferrocyanide  precipitates  cupric  ferrocyanide,  reddish- 
brown,  insoluble  in  acids,  decomposed  by  alkalies;  a  very  delicate  test  for 
copper  (1  to  200,000) ;  forming  in  highly  dilute  solutions  a  reddish 

Solubility  of  Sulphur  Dioxide,  SOj. — Sulphur  dioxide  is  largely 
used  in  the  hydrometallurigcal  methods  of  extracting  copper  from  its 
ores.  Lunge  gives  the  percentage  of  a  saturated  solution  of  sulphur 
dioxide  in  water,  as  follows: 



Temperature,  degrees 


















194  . 




8 . 6  per  cent. 
7.4  per  cent. 
6 . 1  per  cent. 
4.9  per  cent. 
.3 . 7  per  cent. 

2.6  per  cent. 

1 . 7  per  cent. 
0 . 9  per  cent. 
0 . 1  per  cent. 

The  normal  quantity  of  SOj  in  burner-gas  from  brimstone  burners 
is  11.23  per  cent,  by  volume  and  8.75  per  cent,  from  burning  pyrities. 

Sulphur  dioxide  from  roasting  furnaces  is  much  more  dilute;  muflBe 
furnaces  give  a  very  much  more  concentrated  gas  than  reverberatories. 

SOLUBILITY  OF  SO^  IN  WATER  (Watts  Dictionary) 
Absorbed  by  1  grm.  of  water  at  760  mm. 

Temperature,  degrees  C. 



C.c.  SO, 





1        0.065 








j        0.045 


One  liter  of  SO^  weighs  2.86336  grm.  1  cu.  ft.  weighs  0.1787  lb. 
With  water,  sulphur  dioxide  does  not  form  sulphurous  acid  proper, 
HjSOg.  The  sulphur  dioxide  dissolves  pretty  freely  in  water,  and  this 
solution  behaves  in  every  way  as  if  it  contained  the  real  acid,  H2SO3. 

The  solution  of  SO2  by  volume  in  water  at  various  temperatures  is 
as  follows: 

1  volume  of  water  at  0°  C—  32°  F.  dissolves  79.789  volumes  SO^ 
1  volume  of  water  at  20°  C—  68°  F.  dissolves  39.374  volumes  SO^ 
1  volume  of  water  at  40°  C— 104°  F.  dissolves  18.766  volumes  SO, 



Classification  and  General  Consideration. — Hydrometallurgical  proc- 
esses for  the  extraction  of  copper  from  its  ores  or  matte  may  be  con- 
sidered as: 

Purely  Chemical  and 
In  the  purely  chemical  processes  the  copper  is  dissolved  and  precip- 
itated by  chemical  reagents;  in  the  electrolytic  processes,  the  copper  is 
dissolved  chemically  but  the  precipitation  is  effected  electrolytically, 
accompanied,  usually,  by  regeneration  of  the  solvent. 

Chemical  Processes. — These  may  be  classified  as  follows,  based  mostly 
on  the  solvent  employed: 

Alkali  processes, 
Sulphite  processes. 
Sulphate  processes. 
Chloride  processes. 

Nitric  acid,  by  means  of  which  the  copper  would  be  dissolved  as  the 
nitrate  or  sulphate,  has  been  frequently  suggested  as  a  solvent  of  copper 
from  its  ores.  The  fixation  of  atmospheric  nitrogen  by  electricity,  offers 
a  cheap  way  of  producing  nitric  acid  at  the  mines.  There  are,  however, 
inherent  difficulties  to  the  use  of  nitric  acid  which  makes  its  application 
questionable.  Nitric  acid  is  the  best  known  solvent  of  copper,  but  it  is 
also  an  excellent  solvent  of  all  the  impurities  in  the  ore,  so  that  insur- 
mountable difficulties  may  be  expected,  both  in  the  solution  of  the  copper 
and  in  its  precipitation,  if  regeneration  of  the  solvent  is  desired. 

The  applicability  of  any  solvent  process  to  the  extraction  of  copper, 
depends  fundamentally  on  the  character  of  the  ore.  All  acids,  likely  to 
be  used  in  a  solvent  process,  react  more  or  less  with  other  elements; 
when  so  consumed  the  acids  are  unavailable  for  useful  work,  and  fre- 
quently bring  into  solution  ingredients  which  are  positively  harmful 
The  elements  most  detrimental  to  acid  processes  are: 




To  these  may  be  added  iron,  arsenic,  antimony,  bismuth;  but  these  ele- 
ments need  not  necessarily  be  fatal  to  an  acid  process,  no  matter  in  what 
proportion  they  occur  in  the  ore. 

If  lime,  magnesia,  zinc  or  manganese  occur  in  the  ore  in  large  quanti- 
ties, acid  processes  are  not  applicable.  What  the  limit  is,  can  only  be 
determined  by  direct  experiment.  Chemical  analysis  of  the  ore,  while 
instructive,  cannot  be  relied  upon  to  determine  the  applicability  of  an 
acid  process.  Lime,  for  example,  is  only  detrimental  in  certain  combina- 
tions, as  the  oxide  or  carbonate.  In  many  ores  where  sulphuric  acid  has 
been  a  factor  in  the  deposition  or  in  the  oxidation  of  the  vein  matter, 
much  of  the  calcium  will  be  found  as  sulphate,  which  is  not  particularly 
injurious  either  in  a  sulphate,  sulphite  or  chloride  process.  All  of  this 
calcium,  however,  would  usually  be  estimated  as  lime  (CaO),  although  it 
occurs  as  sulphate  (CaSOJ.  Magnesia  is  highly  injurious  as  oxide  and 
carbonate,  but  magnesia  is  not  as  widely  distributed  as  lime,  in  injurious 

Alumina  is  widely  distributed,  but  its  presence,  while  undesirable,  is 
not  necessarily  particularly  injurious.  Much  depends  on  its  mineralogical 
combinations.  In  Cripple  Creek,  sulphuric  acid  has  been  used  for  many 
years  in  connection  with  the  chlorination  of  those  ores,  which  contain 
from  15  to  20  per  cent,  alumina.  Zinc,  especially  as  the  oxide,  is  in- 
jurious, because  it  is  readily  soluble  in  acids,  and  as  yet  no  practicable 
method  has  been  found  for  its  economic-  precipitation.  Electrolytic 
precipitation  offers  a  plausible  way  of  recovering  the  zinc,  and  is  in 
practical  use  in  several  plants,  but  its  general  adoption  is  by  no  means 

Many  oxidized  ores  are  improved  by  roasting.  All  sulphide  ores,  with 
the  possible  exception  of  certain  chalcocite  deposits,  should  be  roasted 
before  chemical  treatment,  no  matter  what  the  nature  of  the  chemical 
treatment  may  be. 

The  treatment  of  raw  sulphide  ores  has  never  met  with  much  encour- 
agement, and  the  cause  for  this  is  reasonable  enough.  The  highly  oxi- 
dized ore  is  in  the  best  possible  condition  for  the  application  of  any 
solvent,  and  it  is  difficult  to  conceive  of  any  oxidizing  process,  or  sub- 
stitution for  an  oxidizing  process,  cheaper  and  more  satisfactory  than 
roasting.  It  is  true  that  low  grade  ores  have  been  treated  in  Spain  and 
Portugal  by  natural  weathering,  and  for  a  while  with  ferric  chloride  or 
ferric  sulphate,  but  the  use  of  ferric  chloride  has  long  since  been  aban- 
doned, and  the  slow  process  of  weathering,  in  which  years  are  required 
to  get  an  adequate  extraction,  is  perhaps  nowhere  else  applicable. 

It  may  be  considered,  therefore,  in  hydrometallurgical  processes, 
that  the  application  of  the  solvent  is  to  the  oxidized  ores. 

Many  of  the  metal  compounds  as  found  either  in  raw  or  roasted  ore, 
have  the  faculty  of  reducing  the  ferric  to  the  ferrous  salts,  the  respective 


metals  being  thereby  brought  into  solution.  This  is  notably  the  case 
with  ferric  sulphate,  FejCSOJ  3,  and  ferric  chloride,  FeClj.  Both  of  these 
substances  have  been  used,  and  have  been  extensively  experimented  with 
in  the  reduction  of  copper  ores,  by  hydrometallurgical  methods.  The 
leaching  method  at  present  used  at  Rio  Tinto,  Spain,  is  based  mostly 
on  the  solvent  action  of  ferric  sulphate,  and  the  Doetsch  process,  formerly 
used  there  extensively,  was  based  on  the  solvent  action  of  ferric  chloride. 




The  alkali  processes  have  not  met  with  much  encouragement  in  the 
hydrometallurgical  extraction  of  copper  from  its  ores.  This  is  due  largely 
to  the  slow  and  low  solubility  of  copper  in  a  solution  of  the  alkalies. 
Ammonia  and  ammonium  compounds  are  about  the  only  alkaline  solvents 
which  have  been  tried  on  a  commercial  scale.  The  oxides  and  carbon- 
ates of  copper  are  quite  readily  soluble  in  ammonia  but  the  solution  should 
take  place  in  tight  receptacles  as  the  volatility  of  the  gas  in  aqueous 
solution  is  quite  perceptible.  If  ammonium  carbonate  is  used  on  cal- 
careous ores  there  should  be  no  sulphates  present,  because  they  would 
be  decomposed  into  ammonium  sulphate  and  calcium  carbonate. 

After  the  copper  is  dissolved  the  ammoniacal  copper  solution  is 
boiled,  and  the  black  oxide  of  copper  precipitated.  The  ammonia  vapor 
boiled  off  may  be  condensed  in  towers  and  used  on  another  charge  of  ore. 
The  recovery  of  the  ammonia  from  the  salts  formed  in  the  boiled-out 
solution  may  be  accomplished  by  means  of  lime  and  steam. 

According  to  Schnabel  "■  experiments  hitherto  tried  in  using  ammonia  or 
ammonium  carbonate  have  failed  because  ammonia-tight  vessels  were  not 
employed,  and  because  the  precipitation  of  the  copper,  for  which  iron 
cannot  be  employed,  was  performed  by  means  of  hydrogen  sulphide, 
calcium  sulphide,  or  barium  sulphide.  By  the  use  of  iron  vessels,  how- 
ever, loss  of  ammonia  may  be  avoided,  but  such  apparatus  has  proved 
complicated  and  expensive  to  operate.  The  oxides  of  zinc,  nickel,  cobalt, 
etc.,  are  also  soluble  in  ammonia  or  ammonium  carbonate. 

Sodium  carbonate  has  been  suggested  as  a  solvent  of  copper  from 
oxide  and  carbonate  ores,  but  the  solubility  of  copper  in  sodium  carbon- 
ate is  so  unsatisfactory  that  experiments  along  these  lines  have  not  been 
encouraging.  Copper  is  also  slightly  soluble  in  concentrated  solutions 
of  the  caustic  alkalies.  Schneider  ^  purposes  increasing  this  solubility  by 
the  addition  of  glycerin. 

The  Mosher -Ludlow  Ammonia-cyanide  Process.' — This  process  depends 
upon  the  principle  that  ammonia,  NH3,  at  the  ordinary  temperature 
forms  soluble,  stable  compounds  with  the  oxides,  hj'droxides  or  carbon- 
ates of  copper,  zinc,  nickel,  or  cobalt,  such  as  Cu(NH3)2. 

'Handbook  of  Metallurgy,  Vol.' I,  p.  204. 
^U.  S.  patent  932,  643,  Aug.  31,  1909. 
'Electrochemical  and  Metallurgical  Industry,  March  1908. 




These  ammonia  metal  compounds  are  readily  dissolved  by  water 
containing  a  small  excess  of  ammonia  over  that  required  to  form  the 
soluble  compound.     This  is  the  leaching  step  of  the  process. 

The  step  of  precipitation  depends  on  the  fact  that  those  soluble 
ammonia-metal  compounds  break  up  with  great  ease  at  the  boiling 
point  of  water  into  the  oxide  or  hydrate  of  the  metal,  which  almost 
instantly  settles  as  a  heavy  precipitate,  while  the  ammonia,  originally 
combined,  is  set  free  to  be  reabsorbed  in  cold  water  or  boiled-out  solution 
for  use  over  and  over  again. 

Fig.  44. — Moser-Ludlow  ammonia-cyanide  process.^     Diagrammatic  sketch. 

Where  the  percentage  of  copper  is  large  the  aim  is  to  first  extract  as 
much  of  the  copper  as  possible  by  plain  ammonia,  and  to  leave  the  gold 
and  silver  values  to  be  subsequently  extracted  with  a  weaker  ammonia 
solution  containing  fractional  percentages  of  potassium  cyanide.  But 
instead  of  working  it  in  this  way  it  may  be  preferable  in  many  instances 
to  add  the  cyanide  at  once  to  the  ammonia  and  to  simultaneously  ex- 


tract  all  the  values,  including  copper,  gold,  and  silver,  with  an  ammoni- 
acal  solution  containing  one  to  several  pounds  of  cyanide  per  ton.  The 
object  aimed  at  is  to  reduce  the  comsumption  of  cyanide  to  a  minimum 
in  the  presence  of  copper,  thereby  permitting  the  minute  amounts  of 
potassium  cyanide  added  to  the  ammonia  solution  to  simultaneously 
extract  the  gold  and  silver  values. 

To  recover  the  metallic  values  from  the  ammonia-cyanide  copper-gold- 
silver  bearing  solution,  it  is  passed  through  a  continuous  boiling-out  still, 
to  precipitate  the  copper  as  CuO.  The  boiled-out  solution  holding  the 
gold  and  silver  values  is  agitated  with  the  least  amount  of  zinc  dust,  or 
passed  through  zinc  boxes  to  recover  such  gold  and  silver  as  the  boiled- 
out  solution  may  contain. 

For  treating  the  ore  containing  considerable  percentages  of  ammonia- 
soluble  metals,  the  chlorination  barrel,  without  lead  lining  is  recommended. 
In  the  treatment  of  slimes,  agitation  in  a  closed  conical  tank,  with  subse- 
quent filtering  with  any  of  the  well-known  filters. 

On  account  of  the  powerful  oxidizing  action  of  a  solution  of  cupric 
oxide  (CuO)  in  ammonia,  unoxidized  silver  minerals  may  be  attacked  and 
finally  dissolved  in  the  raw  state,  obtaining  in  this  manner  a  percentage  . 
of  extraction  out  of  this  character  of  mill  product  entirely  impossible  by 
ordinary  cyanide  methods. 

It  is  of  the  greatest  importance  to  have  a  thoroughly  ammonia-tight 
equipment,  especially  in  the  second  half  of  the  process,  which  comprises 
the  boiling  out  of  the  ammonia  solution.  This  apparatus  is  designed  so 
that  the  boiling  out  is  carried  on  continuously,  and  it  may  in  some  way  be 
compared  to  the  artificial  ice  and  cold  storage  apparatus  in  which  the 
ammonia  water  is  boiled  in  a  stUl,  the  ammonia  gas  distilled  off,  liquefied 
under  pressure  by  powerful  pumps,  then  permitted  to  expand,  by  which 
the  cooling  effect  is  produced,  and  finally  reabsorbed  in  cold  water  to 
commence  the  same  cycle  of  action  over  and  over  again. 

In  the  Mosher-Ludlow  continuous  boiling-out  apparatus  the  in- 
coming cold  ammonia-copper  solution  is  brought  in  contact  with  the 
heat  of  the  ammonia-steam  vapor  in  the  cooler  and  condenser  and  in  the 
heat  exchanger  on  its  way  to  the  boiling-out  still,  with  the  overflowing 
boiled-out  solution.  In  this  way  an  important  part  of  the  heat  applied 
to  boil  out  the  ammonia  is  passed  to  the  incoming  solution,  thus  saving 
steam  and  fuel. 

The  ammonia-copper  solution  is  taken  from  the  copper-solution  tank 
and  pumped  through  the  inner  coil  of  a  double-pipe  counter-current 
cooler;  thence  it  passes  to  the  heat  exchanger,  thence  to  the  boiling-out 
still,  thence  into  the  exchanger,  where  part  of  its  heat  is  passed  to  the 
incoming  solution;  thence  to  the  settling  tanks,  where  the  precipitate, 
which  is  almost  pure  copper  oxide,  is  allowed  to  settle,  the  liquid  being 
drawn  off  into  the  sump  tanks.     From  the  sunip  tanks  the  boiled-out 


solution  is  pumped  through  a  cooler  up  to  the  boiled-out  solution  tank. 
From  there  it  is  used  as  needed  to  make  fresh  ammonia  solution,  and  as 
wash  water. 

The  ammonia  and  steam  vapor  from  the  boiling-out  still  passes  up  to 
the  cooler  and  condenser  into  the  annular  spaces  between  the  two  pipes, 
parting  with  a  fraction  of  its  heat  to  the  incoming  ammonia-copper 
solution  in  the  inner  pipe.  From  the  cooler  and  condenser  the  ammonia 
water  flows  into  the  ammonia  solution  tanks  where  it  is  diluted  and  is 
ready  to  be  sent  to  the  leaching  system.  A  supply  of  cold  water  is  kept 
flowing  over  the  outside  of  the  cooler  and  condenser  and  the  boiled-out 
solution  cooler.  The  precipitate  is  removed  from  the  settling  tank  from 
time  to  time,  as  required. 

It  is  estimated  that  the  heat  required  in  boiling  out  2  per  cent,  ammo- 
nia-copper solution  on  a  100-ton  (24-hour)  basis,  and  precipitating  the 
copper  as  black  oxide  =79.9  per  cent,  copper,  is  51,083,000  b.  t.  u.  of 
which,  however,  according  to  experience,  60  per  cent,  is  saved  by  the 
heat  exchanger,  so  that  the  net  heat  required  amounts  to  34,333,000 
b.  t.  u.  In  practice  this  required  heat  can  be  obtained  from  1  1/2  to  2 
tons  of  good  coal,  and  with  a  boiler  of  from  35  to  40  h.  p. 


A  solution  of  sulphur  dioxide  in  water  may  be  regarded  as  sulphur- 
ous acid,  (HjSOg).  Copper  oxide  and  carbonate  are  soluble  in  sulphurous 
acid,  the  sulphide  is  not. 

Various  methods  have  been  suggested  for  the  practical  use  of  sulphur- 
ous acid  as  a  solvent  of  copper  from  its  ores.  All  of  these  refer  more  di- 
rectly to  the  precipitation  of  the  copper  from  the  sulphurous  acid 
solution,  rather  than  the  solution  of  the  copper  from  the  ore.  Copper 
sulphite  is  not  soluble  in  water,  but  is  readily  soluble  in  excess  of  sulp- 
hurous acid.  Copper  sulphite  is  an  unstable  salt,  which  is  slowly 
changed  into  a  mixture  of  cupro-cupric  sulphite  and  cupric  sulphate,  as 
shown  by  the  following  reactions : 

(1)  3CuO-F3S02  =  3CuS03. 

(2)  3CUSO3  +  CuO  =  Cu2S03,CuS03  +  CuSO ,. 

The  cupro-cupric  sulphite  is  only  very  slightly  soluble  in  water,  but 
is  quite  soluble  in  sulphurous  acid  or  in  a  solution  of  copper  sulphate. 

Neill  Process. — This  process  consists,  first  in  subjecting  the  ore  to  the 
action  of  sulphurous  acid  to  dissolve  the  copper,  and  second,  in  heating 
the  solution  to  drive  off  the  excess  of  sulphurous  acid  and  precipitate  the 
copper  as  sulphite. 

If  the  ore  to  be  treated  is  a  sulphide  it  has  to  be  roasted;  if  the  ore  is 
an  oxide  or  carbonate  it  may  be  treated  without  roasting.  The  sulphur 
dioxide  used  in  the  process  may  be  obtained  from  burning  sulphur  or 


from  roasting  sulphide  ores,  which  if  they  are  copper  ores,  may  be  subse- 
quently treated  in  the  oxidized  condition  with  the  sulphur  dioxide  obtained 
from  the  roasting. 

There  is  less  sulphur  required  to  extract  the  copper  as  sulphite  than  as 
sulphate.  It  takes  98  lb.  of  sulphuric  acid  to  dissolve  63  lb.  of  copper,  as 
sulphate,  while  it  only  takes  32  lb.  of  sulphur,  in  the  form  of  sul- 
phurous acid,  to  dissolve  the  same  amount  of  copper — 63  lb.  The  sul- 
phurous acid  may  be  applied  either  as  a  gas  or  in  solution,  or  a  combi- 
nation of  both. 

In  practice,  the  ore  is  crushed  to  a  suitable  degree  of  fineness  and 
charged  into  agitation  tanks,  and  agitating  the  ore  in  contact  with  the 
sulphurous  acid;  or  the  agitation  may  be  effected  in  tanks  having  a  conical 
bottom,  by  forcing  sulphurous  acid  mixed  with  more  or  less  air,  through  the 
pulp.  In  this  way  the  sulphurous  acid  is  applied  to  the  ore  while  air 
does  the  agitating.  The  tanks,  or  agitators,  in  which  the  ore  is  treated 
may  be  closed  at  the  top,  so  that  the  excess  of  gas  issuing  from  the  first 
tank  may  be  passed  through  the  second  and  so  on  until  the  sulphur 
dioxide  is  entirely  consumed. 

The  clear  solution,  after  being  separated  from  the  ore  by  either  filtra- 
tion or  decantation,  is  heated  to  a  temperature  sufficient  to  drive  out  the 
excess  of  sulphurous  acid,  thereby  precipitating  the  copper  in  the  form 
of  a  bright-red  and  very  heavy  powder  (cupro-cupric  sulphite).  The 
cupro-cupric  sulphite  settles  at  once  to  the  bottom  of  the  precipitating 
tank,  and  the  supernatant  liquor  may  be  decanted  or  siphoned  from  it, 
or  the  sulphite  may  be  recovered  from  the  solution  by  filtration.  The 
excess  of  sulphurous  acid,  driven  out  of  the  solution  by  heating,  may  be 
recovered  for  re-use.  The  cupro-cupric  sulphite  after  it  has  been  removed 
from  the  tanks,  filtered  and  dried,  contains  about  50  per  cent,  metallic 
copper.  This  precipitate  may  be  heated  in  an  oxidizing  atmosphere  in 
a  furnace  and  the  cupric  oxide  produced,  or  it  may  be  melted  in  a  reducing 
atmosphere,  producing  cuprous  sulphide,  which  m%y  then  be  reduced  to 
metallic  copper  by  the  ordinary  converter  process. 

With  ores  suited  to  the  process,  the  copper  will  pass  into  solution  in 
from  1  to  4  hours.  When  the  copper  is  dissolved  by  sulphurous  acid, 
only  very  small  amounts  of  other  metals  are  dissolved,  and  the  ultimate 
product  is  a  very  pure  copper. 

Should  any  copper  exist  in  the  solution  as  sulphate,  due  either  to 
improper  roasting,  or  of  sulphur  trioxide  in  the  sulphurous  acid,  this 
sulphate  of  copper  will  not  be  precipitated  by  boiling  the  solution,  but 
must  be  precipitated  in  some  other  way.  This  may  be  done  either  with 
iron  or  by  electrolysis. 

If  a  solution  of  cupro-cupric  sulphite  is  heated  at  a  high  temperature 
(about  200°  C.)  and  subjected  to  a  pressure  of  about  25  lb.,  sulphur  diox- 
ide is  liberated,  and  there  is  formed  copper  sulphate  and  metallic  copper. 


About  one-half  of  the  copper  may  in  this  way  be  precipitated  in  the 
metallic  condition. 

Neill  Process  at  Coconino,  Arizona'  According  to  Jennings,  the  orig- 
inal method  as  carried  out  at  Coconino,  consisted  in  treating  the  ore, 
with  sulphur  dioxide,  in  a  series  of  upright  tanks  8  ft.  in  diameter  and 
18  ft.  high.  About  5  tons  of  ore  were  introduced  into  the  tank  half  filled 
with  water,  and  gas  was  forced,  by  means  of  a  compressor,  into  this  mix- 
ture of  ore  and  water,  the  excess  gas  passing  from  the  first  tank  to  a 
similar  one,  also  charged  with  ore  and  water,  and  thence  to  a  third  tank, 
where  it  was  supposed  the  absorption  would  be  complete.  The  gas  was 
not  all  used  up  owing  to  the  difficulty  of  absorbing  sulphur  dioxide  in 
water  when  mixed  with  large  volumes  of  air.  When  the  ore  in  the  lower 
tank  was  leached,  an  operation  which  usually  took  10  hours'  time,  the 
solution  and  the  leached  ore  together  were  dropped  into  a  pressure  tank 
and  thence  passed  into  a  large  filter  press.  The  filter  press  was  a  con- 
stant source  of  trouble  as  it  was  impossible  to  find  a  material  for  the 
filters  which  would  stand  any  length  of  time. 

The  solution  from  the  filter  press  was  heated  by  waste  steam  from  the 
crushing  plant  and  60  per  cent,  of  the  copper  precipitated  as  cupro-cupric 

Jennings  gives  the  weak  points  of  the  Neill  process  as  carried  out  at 
Coconino,  as:  1.  The  attempt  to  saturate  the  water  by  simply  blowing 
the  gas  through  it;  2.  the  poor  agitation  obtained  and  the  consequent 
length  of  time  required  to  leach  a  comparatively  small  amount  of  ore; 
3,  the  dilute  solution  obtained,  1  per  cent,  being  the  maximum  amount 
of  copper  which  can  be  held  in  solution  by  an  excess  of  SOj;  4.  the  ease 
with  which  the  copper  separates  from  these  solutions,  both  in  the  leaching 
tanks,  the  pressure  tank  and  the  filter  tanks,  forming  the  cupro-cupric 
precipitate  throughout  the  mass  of  leached  ore,  and  which  it  was  impos- 
sible to  redissolve  with  sulphurous  acid,  and  5.  the  difficulty  of  treating 
the  remaining  40  per  cent,  of  the  copper  in  solution  as  sulphate  after  the 
60  per  cent,  has  been  precipitated.     Scrap  iron  was  not  cheaply  available. 

It  is  evident  that  many  of  the  weak  points  here  enumerated  by 
Jennings  should  not  present  much  difficulty  in  a  well  designed  plant. 
The  effective  absorption  of  gases  in  liquids,  has  long  been  satisfactorily 
accomplished  in  the  chemical  industry,  by  some  means  of  subdivision; 
agitation  is  accomplished  successfully  on  an  enormous  scale  in  the  cyanide 
process,  where  it  is  necessary  to  bring  air  in  contact  with  the  ore  and 
cyanide  solution,  to  effect  extraction.     In  experiments  made  by  Jennings^ 
on  the  same,  or  similar  copper-bearing  triassic  sandstones  of  northern 
Arizona,  he  succeeded  in  getting  an  extraction  of  9.5  per  cent,  of-  the 
copper,  with  sulphur  dioxide,  by  leaching  the  ore  4  hours.     The  low 
copper  content  of  the  solutions  was  due  to  the  small  excess  of  sulphur 
'E.  P.  Jennings,  E.  and  M.  J.,  Jan.  18,  1908. 
'M.  and  J.,  March  30,  1901. 


dioxide.  In  later  experiments^  Mr.  Jennings  succeeded  in  getting  as 
high  as  2  per  cent,  copper  in  the  solutions. 

It  is  evident  that,  so  far  as  the  operations  at  Coconino  are  concerned, 
it  leaves  the  process  where  it  was  before  the  plant  was  erected.  It 
demonstrated  neither  technical  failure  nor  success. 

Neill,  the  inventor  of  the  process,  in  experiments  carried  out  in  Salt 
Lake  City"  succeeded  in  getting  a  complete  extraction  of  Coconino  ore, 
having  11  per  cent,  copper,  in  6  hours.  He  readily  obtained  a  solution 
carrying  2.5  per  cent,  copper,  and  so  long  as  the  solution  remained  cool 
no  difficulty  was  experienced  with  the  copper  separating  out  in  the  sands, 
or  in  washing  the  sands. 

An  experimental  plant,  installed  at  the  smelter  of  the  Montana  Ore 
Purchasing  Company  at  Butte,  by  Neill,  gave  interesting  results.  The 
material  was  roasted  to  about  2  per  cent,  sulphur  content.  It  was  then 
placed  in  a  wooden  barrel  6  ft.  in  diameter  and  12  ft.  long,  and  SOj  gas 
from  the  roasting  furnace  was  blown  through  the  hollow  trunions  and 
brought  into  more  intimate  contact  with  the  pulp  by  means  of  wooden 
paddles  arranged  on  the  sides  and  periphery  of  the  barrel.  Two  tons  of 
roasted  ore  were  charged  with  5  tons  of  water,  and  after  passing  the 
dilute  roasting  gases  through  the  barrel  for  6  or  8  hours  the  copper  was 
successfully  extracted.  The  barrel  was  dumped  into  a  settling  tank,  the 
solution  drawn  off  by  percolation,  and  the  sands  washed  in  the  same  way. 
This  washing  was  difficult  on  account  of  the  ferric  oxide,  which  being 
flocciilent,  remained  in  suspension  and  formed  a  layer  upon  the  top  of 
the  sands  which  it  was  difficult  to  percolate.  The  sands  after  this  incom- 
plete wash  averaged  0.8  per  cent,  copper,  but  average  samples  taken  from 
the  tank  and  washed  by  agitation  and  decantation,  gave  final  tails  of 
0.31  per  cent,  copper.  The  heads  averaged  3.15  per  cent.,  showing  an 
extraction  of  90  per  cent.  The  solution  was  heated  in  a  wooden  tank  by 
a  steam  coil,  and  the  precipitates,  which  were  slightly  contaminated  with 
alumina,  on  account  of  the  poor  filtration,  amounted  to  64  per  cent,  of  the 
copper  extracted.  The  remainder  was  precipitated  upon  iron  in  a  splash 
tank  and  the  final  solution  turned  to  waste  carried  only  traces  of  copper. 

The  amount  of  iron  consumed  was  exceedingly  small  and  the  reaction 
very  quick,  owing  to  the  fact  that 'the  solution  came  from  the  steam 
tanks  at  nearly  the  boiling  point.     • 

There  were  35  tons  of  material  treated,  and  the  figured  cost  of  the 
operation  compared  favorably  with  anything  being  done  at  that  time  or 
now  in  the  Butte  district.  The  process  was  not  adopted  because  at  that 
time  the  silver  content  of  the  company's  ore  was  high  and  could  not  be 
saved  by  this  method,  and  the  space  necessary  for  the  plant  was  not 

'E.  and  M.  J.,  April  18,  1908. 
'E.  and  M.  J.,  March  14,  1908. 


Van  Arsdale  Process.' — The  van  Arsdale  process  consists  in  precipitat- 
ing copper  from  cupric  sulphate  solutions  and  simultaneously  producing 
sulphuric  acid,  by  adding  to  solutions  of  cupric  sulphate,  sulphur  dioxide 
and  heating  with  or  without  pressure.  The  copper  is  thrown  down  in 
solid  form  which  may  be  subsequently  treated,  while  the  regenerated 
acid  is  applied  to  the  ore  for  the  extraction  of  mor^  copper. 

The  ore  must  contain  the  copper  as  oxide  or  carbonate.  The  original 
solution  for  leaching  must  contain  cupric  sulphate  and  should  contain 
ferrous  sulphate.  The  precipitation  will  proceed  with  cupric  sulphate 
alone,  resulting  in  the  formation  of  a  salt  of  copper  and  the  formation  of 
sulphuric  acid;  by  the  addition  of  ferrous  sulphate  the  reaction  proceeds 
better,  and  by  the  use  of  a  proper  proportion  of  ferrous  sulphate  there  is 
obtained  a  precipitate  of  metallic  copper.  In  practice  the  solutions 
will  always  contain  more  or  less  ferrous  sulphate  dissolved  from  the  ores 
treated.  Good  results  are  obtained  from  a  solution  containing  approxi- 
mately 10  per  cent,  each  of  cupric  sulphate  and  ferrous  sulphate.  To 
precipitate  the  copper,  sulphur  dioxide  is  applied  to  the  solution  to 
nearly  saturation.  The  solution  is  then  heated  to  nearly  the  boiling 
temperature,  whereupon  a  reaction  takes  place,  resulting  in  the  precipita- 
tion of  a  part  of  the  copper  contained  in  the  solution  either  as  metallic 
copper  or  compounds  of  copper,  or  both,  together  with  the  formation  of 
sulphuric  acid.  The  amount  of  copper  precipitated  will  vary  with  the 
composition  of  the  solution  and  also  with  the  pressure  under  which  it 
is  heated.  When  the  solution  is  heated  under  pressure  an  increased 
precipitation  of  copper  is  obtained,  amounting  to  about  50  per  cent,  of 
the  copper  originally  present,  and  an  amount  of  sulphuric  acid  regener- 
ated amounting  to  about  double  that  necessary  to  redissolve  from  ore 
the  amount  of  copper  precipitated.  The  process  being  cyclic,  there  is 
no  particular  harm  in  returning  to  the  ore  a  solution  containing  a  con- 
siderable amount  of  unprecipitated  copper  sulphate. 

The  chemical  reactions  involved  in  the  process  may  be  given  as: 

(1)  3CuS04-F3S02  +  4H20  =  CuS03,Cu2S03+4H2SO,. 
(2)  Cu,S03,CuS03  +  4H2SO,  =  Cu-t-2CuSO,  +  2H2SO,+2S02-F2H,0. 

At  atmospheric  pressure  and  in  the  cold,  a  solution  of  cupric  sulphate 
saturated  with  sulphur  dioxide,  after  standing  for  some  time,  deposits  a 
small  amount  of  cupro-cupric  sulphite.  On  heating  such  a  solution  to 
boiling  and  at  the  same  time  passing  sulphur  dioxide  through  it,  a  larger 
amount  of  copper  is  precipitated,  resulting  finally  in  a  precipitate  of 
metallic  copper,  according  to  equation  No.  2.  When  the  solution,  how- 
ever, is  saturated  with  sulphur  dioxide  and  placed  in  a  closed  vessel  and 
heated  under  pressure,  the  yield  of  precipitated  copper  is  increased  to 
'E.  and  M.J.,  June,  1903;  U.  S.  Patent,  March  31,  1903,  No.  723,949. 


40  and  50  per  cent,  of  the  copper  originally  present,  and  free  acid  is 
formed  according  to  the  above  equations. 

The  degree  of  heat  and  pressure  required  for  the  second  operation 
are  not  high,  it  being  only  necessary  to  heat  the  saturated  solution  to 
nearly  100°  C,  the  pressure  produced  being  about  30  lb.  to  the  square 
inch.     A  lead  lined  steel  tank  may  be  used  for  this  purpose. 

Jumau  found'  the  following  relations  of  temperature  and  proportion 
of  copper  precipitated  from  a  copper  sulphate  solution  saturated  with 
sulphurous  acid,  having  originally  25   grm.  of  copper  sulphate  per  liter. 

Temperatures  Copper  precipitated 
140°  C.  47  per  cent. 

155°  C,  62  per  cent. 

167°  C,  65  per  cent. 

190°  C,  79  per  cent. 


Ordinarily  only  oxidized  ores  such  as  the  oxides  and  carbonates,  can 
be  treated  so  as  to  dissolve  the  copper  as  sulphate.  Sulphides  are  not 
usually  amenable,  practically,  to  direct  treatment.  Roasting  is  desir- 
able. After  the  ore  is  roasted  the  copper  should  be  in  the  form  of  oxide 
and  sulphate,  although  in  improperly  roasted  ore  sulphides  may  still  be 
present.  The  roasted  material  is  then  treated  the  same  as  naturally  oc- 
curring oxidized  ores.     The  copper  may  be  dissolved  as  sulphate  either  by 

Sulphuric  acid,  or 
Metal  sulphates, 

such  as  the  ferric  sulphate.  The  acid,  however,  is  the  solvent  most  uni- 
versally employed. 

If  iron  is  used  as  the  precipitant,  as  it  usually  is,  the  sulphuric  acid 
process  consists  essentially  of  applying  dilute  sulphuric  acid  to  the  oxi- 
dized ores  of  copper,  which  reacts  with  the  oxide  of  copper  as  follows: 

CuO-|-H2S04  =  CuS04-hH20. 

The  copper  sulphate  thus  formed  is  filtered  from  the  ore  and  precipitated 
with  iron,  thus: 

CuSO,-f-Fe  =  Cu+FeSO„ 

the  iron  and  the  copper  changing  places.  The  copper  is  precipitated 
while  the  iron  goes  into  solution  as  ferrous  sulphate.  From  this  reaction 
it  will  be  seen  that,  theoretically,  it  takes  98  lb.  of  sulphuric  acid  to 
dissolve  63.6  lb.  of  copper,  and  56  lb.  of  iron  to  precipitate  63.6  lb.  of 
copper;  1.56  lb.  of  sulphuric  acid  to  dissolve  1  lb.  of  copper  and  0.88  lb. 
of  iron  to  precipitate  it.  These  equivalents  are  for  pure  acid  and  pure 
'U.  S.  Patent  930,967,  Aug.  10,  1909. 


iron.  Commercial  sulphuric  acid  always  contains  more  or  less  water  and 
other  impurities,  and  in  iron  also,  there  is  more  or  less  foreign  matter. 
So  it  may  be  safely  assumed  that  it  will  take  1.75  lb.  of  ordinary  commer- 
cial sulphuric  acid  to  dissolve  1  lb.  of  copper  and  0.95  lb.  of  iron  to  precipi- 
tate a  pound  of  copper. 

The  problem  of  the  commercial  treatment  of  ores  by  sulphuric  acid 
is,  however,  much  more  complicated  than  the  simple  process  here  outlined. 
If  the  ore  contains  foreign  elements  attackable  by  sulphuric  acid,  the 
acid  consumption  may  be  excessive,  and  if  the  copper  sulphate  solution 
entering  the  precipitation  tanks  contains  free  acid  or  ferric  sulphate  the 
consumption  of  iron  for  precipitation  will  also  be  excessive. 

To  determine,  therefore,  whether  any  ore  may  be  economically 
treated  by  a  sulphate  process,  it  is  necessary  to  make  a  direct  test  on  the 
ore,  and  thus  ascertain  the  consumption  of  acid.  The  iron  for  pre- 
cipitation may  be  taken,  in  practice,  at  1.5  to  2.0  lb.  of  scrap  iron  per 
pound  of  copper  produced.  These  factors  being  known,  close  approxi- 
mations can  usually  be  made  as  to  the  commercial  applicability  of  the 
process  in  its  simplest  form. 

Many  improvements  on  the  above  simple  process  have  for  their  basis 
the  cheapening  of  the  solution  of  the  copper,  but  most  of  the  improve- 
ments are  based  on  the  precipitation  and  regeneration  of  the  solvent. 
If  iron  is  used  as  the  precipitant  in  the  simple  process  as  outlined,  the 
acid  is  lost,  so  that  new  acid  has  to  be  supplied  to  the  ore  at  every  cycle 
of  solution.  This  being  the  case,  sulphuric  acid  installations  may  be 
made  at  the  mine  using  the  sulphide  ore  for  the  production  of  the  neces- 
sary sulphur  dioxide,  for  the  manufacture  of  acid.  The  ore,  after  roast- 
ing, may  be  treated  with  the  acid  so  produced,  to  extract  the  copper, 
which  is  in  the  form  of  oxide  or  sulphate  and  is  readily  amenable  to  the 

Sulphide  of  copper,  in  ores,  may  be  converted  into  sulphate  by  the 
action  of  ferric  sulphate,  thus: 

xH2SO,  +  Cu2S+2Fe2(SOj3  =  CuSO,  +  4FeSO,-FS+xH2SO„ 
in  which  x  represents  an  indeterminate  amount  of  the  sulphuric  acid. 
Ferric  sulphate  has  not  been  used  independently  for  method  of  extraction 
on  roasted  oxidized  ores,  owing  to  its  slow  action.  In  imperfectly 
roasted  ore  its  presence  may  be  beneficial  by  promoting  the  formation  of 
sulphate  from  the  remaining  undecomposed  sulphides,  should  any  be 

Sulphur  dioxide,  steam,  and  nitrous  fumes  have  been  applied  to  ore 
to  dissolve  the  copper  as  sulphate;  this  amounts,  essentially,  to  the 
manufacture  of  sulphuric  acid  within  the  ore  mass. 

In  the  application  of  sulphuric  acid  for  the  extraction  of  copper, 
cupric  oxide,  azurite,  malachite,  and  arsenate  of  copper  dissolve  readily; 
phosphate  of  copper  with  more  or  less  difficulty.     Cuprite   (cuprous 


oxide,  CujO) ,  is  not  readily  soluble,  but  if  moistened  with  acid  and  left 
exposed  to  the  air  for  some  time,  it  is  transformed  into  the  cupric  oxide, 
and  is  then  readily  soluble. 

According  to  SchnabeP  at  Stadtberg  in  Westphalia  and  at  Linz  on 
the  Rhine,  ores  containing  1  to  2  per  cent,  of  copper  were  sulphated  by 
means  of  sulphur  dioxide  steam  and  nitrous  gases,  and  then  leached. 
At  Stadtberg  the  ores  were  azurite  and  malachite  disseminated  through 
the  quartzose  shist,  at  Linz,  at  the  Stern  works,  copper  carbonates  and 
phosphates.  The  leaching  vessels  were  tanks  of  brickwork  3  ft.  3  in. 
deep.  Above  the  bottom  proper  these  had  a  false  bottom  of  grating, 
made  of  fire  brick  or  other  acid  proof  material,  supported  by  bricks  on 
edge.  The  ores  were  piled  up  on  this  grating,  the  gases  being  conducted 
underneath  it.  The  gases  were  respectively  generated  by  roasting  iron 
pyrites  in  shaft  furnaces  and  zinc  blende  in  muffle  furnaces,  and  by  treat- 
ing Chili  niter  with  sulphuric  acid.  The  sulphur  dioxide,  nitrous  fumes, 
and  steam  together  formed  sulphuric  acid  which  converted  the  copper 
compounds  into  sulphates.  After  8  to  10  days  the  copper  sulphate  was 
dissolved  out  by  means  of  water  or  of  the  acid  mother  liquor  left  after 
precipitating  the  copper.  The  leaching  was  so  conducted  that  fresh  water 
or  the  copper-free  mother  liquor  was  allowed  to  attack  the  most  com- 
pletely exhausted  ore,  while  the  almost  saturated  solution  was  run  on  to 
fresh  ore  until  is  was  fully  saturated  (22  to  26°  B).  The  liquor  that 
drained  away  ran  into  receivers,  whence  it  was  again  pumped  on  to  the 
ores.  This  process,  which  extracted  the  copper  down  to  1/4  per  cent., 
has  long  ago  been  abandoned.  At  Stadtberg  sulphuric  acid  was  re- 
placed by  the  cheaper  hydrochloric  acid  as  long  as  the  supply  of  oxi- 
dized ores  lasted. 

Acid  Plants  at  the  Mine. — Conditions  are  frequently  ideal  for  acid 
manufacture,  for  leaching  purposes,  at  the  mines,  if  suitable  sulphide  ore 
is  available.  If  the  mine  produces  both  sulphide  and  oxide  ores,  the 
sulphides  may  be  roasted  and  the  sulphurous  gases  converted  into  acid 
which  may  then  be  used  to  leach  both  the  oxidized  and  roasted  ores. 
If  the  sulphide  ore  is  low  in  sulphur,  concentration  will  usually  be  neces- 
sary to  get  a  material  sufficiently  high  in  sulphur  to  make  a  gas  suitable 
for  sulphuric  acid  manufacture. 

The  sulphide  ore,  or  concentrates  are  usually  roasted  in  a  furnace  of 
the  McDougal  type,  and  the  sulphur  dioxide  gas  passed  from  the  roasting 
furnace  into  a  series  of  leaden  chambers,  where  coming  in  contact  with 
gaseous  nitric  acid  and  steam  it  becomes  converted  into  sulphuric  acid. 
The  nitric  acid  gas  is  produced  by  the  action  of  sulphurc  acid  and  nitrate 
of  soda,  and  passes  along  with  the  sulphurous  gases  into  the  lead  cham- 
bers. The  combined  gases,  together  with  the  seam  and  air,  mix  in  the 
chambers  and  condense  as  sulphuric  acid.     This  is  known  as  chamber 

'Handbook  of  Metallurgy,  Vol.  I,  p.  200. 


acid  and  has  a  specific  gravity  of  about  52° B.  and  contains  about  65.2 
per  cent.  HjSO^.  The  chamber  acid,  in  commercial  plants,  is  then  con- 
centrated to  66°B.  containing  93.5  per  cent.  H^SO^,  and  this  is  the  ordi- 
nary acid  of  commerce,  or  oil  of  vitriol. 

In  the  manufacture  of  sulphuric  acid,  for  leaching  purposes,  the  process 
will  be  somewhat  cheaper  than  the  manufacture  of  66°  B.  acid  for  com- 
merce, because  no  purification  will  be  required,  and  the  chamber  acid 
may  be  used  without  further  concentration.  On  the  other  hand,  the 
manufacture  of  acid  for  leaching  purposes  will  usually  be  conducted  on  a 
small  scale,  and  hence  the  cost  of  operation,  per  unit  of  acid,  will  be  largely 
increased  over  that  of  the  large  commercial  plants. 

In  the  manufacture  of  sulphuric  acid,  a  chamber  space  cf  from  15  to 
25  cu.  ft.  should  be  provided  per  pound  of  sulphur  burned  in  24  hours,  and 
nitrate  of  soda  will  be  consumed  in  amounts  varying  from  3  to  5  per  cent, 
of  the  sulphur  burned. 

The  cost  of  manufacturing  the  acid  at  the  mine  will  vary  within  wide 
limits.  The  selling  price  of  66°B.  acid,  f.  o.  b.  works,  in  commerical 
plants,  is  about  $18.00  per  ton,  but  this  cannot  be  made  the  basis  of  costs 
the  mine,  for  the  reason  that  such  acid  is  made  under  the  best  possible 
economic  conditions.  On  the  other  hand,  if  acid  is  manufactured  at  the 
mine,  assuming  the  ore  to  be  a  sulphide  or  to  contain  sufficient  sulphides 
for  acid  manufacture,  there  would  be  no  expense  for  roasting  and  no 
expense  for  sulphur,  for  in  any  event,  the  suphides  would  have  to  be 
roasted,  and  the  sulphurous  gases  would  otherwise  be  wasted. 

Sulphuric  Acid  Leaching  of  Oxidized  Copper  Ores,  at  Clifton,  Arizona. 
— The  Arizona  Copper  Company  have  been  leaching  oxidized  surface 
ores  at  Clifton,  on  a  large  scale,  since  1893.  The  following  description 
of  the  work  was  prepared  by  F.  N.  Flynn,  the  Company's  Metallurgist.' 

"Four  groups  of  mines,  in  the  Metcalf  District,  have  contributed  this  class 
of  ore,  but  the  Metcalf  mine  has  furnished  the  principal  part  of  the  tonnage. 
The  occurrence  of  the  ore,  and  the  method  of  mining  have  been  described  by 
Mr.  Peter  B.  Scotland,  in  the  Eng.  and  Min.  Jour.,  July  16,  1910. 

"  The  ores  are  lowered  down  the  hillside  by  means  of  inclined  tramways  to  the 
railroad  bins  at  Metcalf.  Trains  of  40-ton,  bottom  dump  cars  are  hauled  over 
the  Coronado  Railway,  36-in.  gauge  tracks,  to  Clifton,  6.6  miles  distant.  From 
the  railroad  bins,  which  are  common  to  the  various  departments  of  the  works, 
the  ore  is  conveyed  by  a  30-in.  belt  to  the  "Oxide  Mill"  bin,  located  near  the 
smelting  plant. 

"  The  gangue  of  the  ore  partakes  of  the  character  of  the  mine  formations — 
highly  altered  sedimentary  and  igneous  rocks.  The  granite-porphyry,  quartzite, 
shale,  and  limestone  have  all  been  more  or  less  altered,  resulting  in  a  mass  of 
quartz  grains,  kaolin,  serecite,  silicious  hematite,  magnetite,  limonite,  garnet  and 

'Private  communication  from  Norman  Carmichael,  Gen.  Manager,  Arizona  Cop- 
per Co.,  Nov.  1,  1911. 


various  other  silicates  of  alumina.  Fortunately  the  limestone  was  completely 
altered,  and  the  calcium  sulphate  almost  completely  removed. 

"The  copper-bearing  minerals,  in  the  order  of  their  importance,  are:  Mala- 
chite, copper-pitch-ore,  azurite,  impure  chrysocolla,  cuprite,  chalcocite,  chalco- 
pyrite,  native  copper  and  brochantite.  Malachite  occurs  throughout  the  upper 
part  of  the  deposit,  and  is  the  all-important  mineral.  The  impure  chrysocolla  is 
neither  brochantite  nor  pure  chrysocolla.  Specimens  of  this  impure  mineral, 
containing  22  per  cent,  of  copper,  contain,  but  0.05  per  cent,  of  sulphur.  The 
copper  is  readily  soluble  in  very  dilute  sulphuric  acid  without  effervescence. 
Cuprite  occurs  in  the  shale,  usually  unassociated  with  other  copper  minerals, 
except  native  copper  in  small  amounts.  It  occurs  in  very  thin  flakes,  and  in  such 
a  manner  as  to  suggest  its  deposit  direct  from  copper  sulphate  solutions,  the 
latter  having  been  hydrolized  by  absorption.  Chalcocite  is  found  in  the  porphyry, 
partly  and  completely  replacing  the  pyrites. 

"The  milling  ores  vary  between  2.5  and  3.0  per  cent,  copper,  and  analyze 
about  as  follows : 

SiOj,  59 . 0  per  cent. 

AI2O3,  18.0  per  cent. 

Fe,  9 . 0  per  cent. 

Mn,  0 . 1  per  cent. 

CaO,  0 . 1  per  cent. 

MgO,  0 .  05  per  cent. 

S,  1.0  per  cent. 
Au,  Ag,  Pb,  and  Zn,  traces. 

About  315  tons  of  crude  ore  are  treated  per  day. 
The  crushing  plant  consists  of : 

One  10-in.  by  20-in.  jaw  crusher. 

Two  sets  of  12-in.  by  36-in.  rolls. 

One  belt  elevator. 

One  trommel,  with  3/4-in.  and  5/8-in.  holes  when  new. 

Water  is  fed  under  the  first  set  of  rolls. 

The  3/4-in.  oversize  goes  to  the  second  set  of  rolls. 

The  3/4-in.  to  5/8-in.  size  is  finished  product,  also  the  5/8-in.  undersize. 
Nothing  is  reorushed.     The   crushing  plant   and  jigs  run  17  hours  per  day. 

"  Material  between  3/4-in.  and  5/8-in.  goes  to  one  two-compartment  Hartz  jig, 
making  top  concentrates  and  tails.     The  hutch  product  returns  to  the  ore  stream. 

"The  5/8-in.  undersize,  including  the  crushing  water,  goes  to  one  five-ccm- 
partment  Hancock  jig,  making  concentrates  and  tails. 

"The  concentrates  analyze  about  as  follows: 

SiOj,  37.0  per  cent. 

AI2O3  11.0  per  cent. 

Fe,  22 . 0  per  cent. 

Mn,  0.2  per  cent. 

CaO,  0,04  per  cent. 

MgO,  0.1  per  cent. 

S,  5.0  per  cent. 

Cu,  7.0  to  10.0  per  cent. 


"The  [specific  gravity  of  the  minerals  entering  the  concentrate  runs  between 
3. 7(1  and  4.63  on  sizes  larger  than  1/8  in.  Eighty-five  per  cent,  of  the  con- 
centrate is  larger  than  eighth  inch.  From  20  to  2.)  per  cent,  of  the  values  are 
recovered  as  concentrate,  at  a  ratio  of  about  10  to  1. 

"A  belt  elevator  lifts  the  concentrates  to  a  bin  on  the  smelter  charge  floor, 
from  which  they  are  wheeled  direct  to  the  blast  furnaces. 

"The  tails  bin  serves  as  a  dewaterer.  The  slime  water  is  pumped  to  settUng 
ponds  three-quarters  of  a  mile  distant.  The  slime  carries  2.4  per  cent,  copper  and 
too  much  soluble  alumina  to  permit  of  leaching  it  with  the  gravel  in  the  present 
plant.  The  tailings  vary  from  1  in.  (due  to  the  wear  of  the  trommel)  down  to  the 
finest  sands;  75  per  cent,  is  larger  than  eighth  inch.  The  tailings  are  hoisted  by 
inclined  skip  to  a  receiving  bin,  centrally  located  over  the  top  of  the  leaching 

"The  Joy  Mine  at  Morenci  furnishes  the  pyrite  for  the  manufacture  of  sul- 
phuric acid.     The  pyrite  analyses: 

SiOj,  9.0  per  cent. 

AI2O3  4.0  per  cent. 

Fe,  38.0  per  cent. 

S,  38.0  per  cent. 

Zn,  2,0  per  cent. 

Cu,  Variable;  usually  under  1.5  per  cent 

The  pyrite  is  crushed  to  2  in.  The  fines  are  roasted  in  a  Herreshoff  five-deck 
furnace.  The  coarse  material  goes  to  the  lump  burners.  The  cinder  goes  to  the 
blast  furnaces. 

"The  resulting  acid  is  52°  Baume.  The  capacity  of  the  acid  plant  is  10  tons 
per  day. 

"In  the  leaching  department,  small  circular  wooden  tanks  are  used.  There 
are  twelve  ore  tanks  of  13  tons  capacity  each,  or  156  tons  total  capacity.  These 
have  an  inner  lining  of  lighter  staves.  Between  the  outer  and  inner  staves  an 
acid  proof  cement  preparation  is  used.  The  tops  of  the  staves  are  covered  with 
sheet  lead.  A  false  bottom  of  plank,  slightly  inclined  toward  the  center,  and 
perforated  with  hoels.  serves  as  a  filter  bottom. 

"The  ore  (jig  tails)  is  charged  to  the  tank  by  means  of  fixed  launders  over- 
head, connected  with  the  ore  supply  bin  at  the  top  of  the  mill. 

"  The  leached  tailings  are  discharged  through  a  circular  opening  in  the  bottom 
of  the  tank  at  the  center.  The  opening  is  closed  bj*  means  of  a  wooden  plug, 
extending  up  through  the  ore  charge,  and  suspended  from  chain  blocks.  The 
tails  are  flushed  from  the  ore  tanks  and  hoisted  to  the  railroad  bins,  from  which 
they  are  hauled  away  in  50-ton  bottom  dump  cars.  Each  ore  tank  is  provided 
with  a  belted  centrifugal  pump  for  circulating  liquors  from  bottom  to  top  of 
tank,  or  to  any  other  tank  in  the  series. 

"  Ordinary  circular  wooden  tanks  are  used  in  the  precipitating  room.  Two 
square  tanks  are  used  for  final  precipitation.  In  each  of  these  tanks  is  a  revolving 
drum  or  trommel  made  of  cast  copper.  The  circumference  is  perforated  with 
holes  like  a  trommel.  The  ends  are  closed  and  support  the  axles.  A  small  door 
permits  charging  the  trommel  with  small  pieces  of  wrought  iron.  The  drum  is 
submerged  in  the  solution  up  to  the  bearings,  and  revolves  slowly. 


The  department  is  arranged  on  three  floors.  The  top  floor  for  the  ore  tanks, 
the  middle  floor  for  the  precipitating  tanks  and  trommels,  and  the  ground  floor 
for  cement  copper. 

"The  ore  charge  (jig  tails)  analyze: 

SiOj,  60.0  per  cent. 

AI2O3,  16.0  per  cent. 
Fe,  8.0  per  cent. 

Mn,  0. 1  per  cent. 

CaO,  0 .  06  per  cent. 

MgO,  0 .  03  per  cent. 

S,  0.4  per  cent. 

Cu,  2 . 0  to  2 . 6  per  cent. 

"In  practice  three  tanks  are  leached  as  a  unit  as  regards  solutions,  although 
the  tanks  are  charged  and  discharged  singly  and  at  intervals. 

"A  charge  of  the  solution  circulates  in  each  of  the  three  tanks  for  4  hours,  or 
12  hours  total  time,  more  or  less.  In  this  manner  a  charge  of  ore  is  leached  with 
three  strengths  of  liquor,  approximately  as  follows : 

4  hours  with  '  Strong  Liquor.'     High  in  copper  and  low  in  acid. 

4  hours  with  'Weak  Liquor.'     Lower  in  copper  and  higher  in  acid. 

4  hours  with  'Acid  Solution.' 

0 . 5  hours  with  Water. 

0 . 5  hours  Charging  and  Discharging. 

13  hour  cycle. 

"  Sulphuric  acid  is  added  to  the  resulting  wash  water,  to  make  the  'Acid 
Solution,'  which  in  turn  makes  'Weak  Liquor,'  'Strong  Liquor'  and  'Copper 

"The  'Copper  Liquor'  goes  to  the  precipitating  tanks  with  a  very  small 
fraction  of  a  per  cent,  of  free  acid.  Usually  the  free  acid  is  too  small  to 

"The  acid  consumed  in  leaching,  per  pound  of  pure  copper  recovered,  amounts 
to  2.6  lb.  of  52°  B.  acid. 

"The  'Copper  Liquor'  from  the  various  tanks  is  collected  in  a  distributing 
tank,  from  which  it  flows  continuously  in  a  small  regulated  stream  to  the  pre- 
cipitating tanks.  The  precipitating  tanks  are  connected  in  series,  and  use  scrap 
iron  of  all  descriptions — all  of  which  comes  from  the  works. 

"From  the  last  tank  in  the  series  the  liquor,  low  in  copper,  goes  to  the  'trom- 
mel tank'  in  'charges.'  Small  pieces  of  wrought  iron  are  charged  to  the  trom- 
mel. Usually  this  consists  of  ties  from  cotton  bales,  cut  in  strips  of  a  foot  in 

"A  'charge'  of  liquor  requires  from  10  to  30  minutes  in  contact  with  the 
revolving  trommel  to  complete  the  copper  precipitation.  The  finished  charge 
containing  its  cement  copper  is  flushed  to  settling  tanks. 

"The  decanted  liquor,  practically  free  from  copper,  either  in  solution  or  as 
precipitate,  is  passed  through  another  set  of  overflow  tanks,  containing  tin  cans, 
before  going  to  waste. 

"The  resulting  'Iron  Liquor'  is  pumped  to  an  earthen  reservoir,  where  it 
soaks  into  the  ground. 


"  The  cement  copper,  after  settling  to  separate  the  solution,  is  accumulated  in 
the  patio,  and,  when  sufficiently  dry,  is  moulded  into  large  bricks,  by  hand. 
After  sun  drying  for  a  month,  they  are  sufficiently  dry  to  be  fed  to  the  copper 
converters.     The  cement  copper  averages  72  per  cent,  copper. 

"The  plant  is  handled  by  a  superintendent  and  two  white  foremen  with 
Mexican  laborers.  The  foremen  make  frequent  mill  tests  for  copper,  impurities 
and  free  acid.  The  working  strength  of  all  solutions  is  governed  by  Baume 

"No  general  rule  can  be  followed  for  the  strength  of  the  various  solutions, 
because  the  ore  is  quite  variable  in  composition,  both  as  regards  copper  and  other 
soluble  salts.  When  ores  are  met  with  which  show  a  readily  soluble  gangue,  the 
acid  strength  of  the  solution  is  reduced  from  that  used  or  on  more  silicious 

Leaching  Plant  at  the  Snowstorm  Mine. — At  the  Snowstorm  mine, 
Larson,  Idaho,  there  has  been  in  operation  for  some  years  a  leaching 
plant  of  250  tons  daily  capacity.  The  ore  deposit  of  the  Snowstorm 
mine  consists  of  disseminations  of  bornite,  chacolcite,  and  chalcopyrite 
in  certain  beds  of  Revett  quartzite.  The  greater  part  of  the  sulphide 
has  however,  been  oxidized  to  cuprite,  malachite,  and  chrysocolla.  The 
various  prospects  are  on  metasomatic  fissure  veins  carrying  chalcopy- 
rite, chalcocite  or  bornite,  with  quartz,  dolomite,  or  siderite.  In  the 
lower  workings  of  the  mine  the  ore  occurs  as  sulphide,  containing  only  a 
very  small  portion  of  the  copper  in  the  oxidized  condition,  but  no  attempt 
has  been  made  to  leach  the  sulphide  ore. 

The  oxidized  ores  average  from  2  1/2  to  3  1/2  per  cent,  copper,  7  oz. 
in  silver,  and  $1.00  in  gold,  per  ton.  The  ore  is  crushed  and  run  into  three 
agitators,  where  it  is  treated  with  bleaching  powder  and  a  10  per  cent, 
solution  of  sulphuric  acid.  By  this  method  chlorine  is  slowly  released, 
which  chlorinates  the  gold  and  silver,  and  to  some  extent  attacks  the 
small  quantities  of  sulphide  in  the  oxidized  ores. 

The  copper  solution  goes  from  the  agitators  through  a  series  of  six 
settling  tanks,  after  which  it  is  precipitated  with  scrap  iron.  The  resi- 
dues, containing  the  silver  chloride,  are  treated  with  sodium  thiosulphate 
(hyposulphite)  to  dissolve  the  silver,  and  the  solution  so  obtained  is  passed 
through  settling  tanks  and  the  silver  precipitated  from  the  clear  solution 
with  sodium  sulphide.  The  silver  sulphide  precipitate  is  filtered  and 
shipped  for  refinement.  The  process  is  said  to  save  90  per  cent,  of  the 
assay  value  of  the  ore. 

Copper  Leaching  Plant  at  the  Gumeshevsky  Mine,  Russia.' — The 
Sissert  property,  of  which  the  Gumeshevsky  copper  mine  is  a  part,  is  one 
of  the  largest  concessions  in  the  Urals  and  was  originally  obtained  for 
working  iron  ore  deposits.     As  early  as  1727  two  important  copper  de- 

'Inst.  of  Min.  and  Met.  Bull.,  No.  65;  Trans.  I.  M.  M.,  XIX,  212;  Min.  Ind., 
1910,  210. 


posits  were  discovered  on  the  property.  These  were  worked  intermit- 
tently for  more  than  100  years. 

The  Gumeshevesky  mine  was  shut  down  in  1871.  From  old  data 
available  it  appears  that  the  mine  is  a  contact  between  limestone  and 
diorite.  Oxidized  copper  ore  occurs  in  a  clay  formation  along  the  con- 
tact where  there  are  old  workings,  extending  about  2  miles  in  length 
and  possibly  1000  ft.  wide.  The  mine  was  worked  for  the  oxidized  ores 
only.  The  deepest  shaft  is  500  ft.  The  material  raised  from  the  shaft 
was  evidently  hand  picked  and  only  the  large  lumps  of  oxidized  ore 

The  dump  consists  of  clay  material  with  the  fine  ore  that  escaped 
hand  picking.  It  covers  an  area  of  about  20  acres,  with  an  average 
depth  of  17  ft.  This  was  thoroughly  sampled  and  estimated  to  contain 
531,000  cu.  yd.  of  material  carrying  23  lb.  of  copper  per  cubic  yard  (about 
0.79  per  cent). 

In  brief,  the  process  of  treatment  consists  of  leaching  the  crushed 
material  with  dilute  sulphuric  acid,  then  precipitating  the  copper  from 
solution  with  pig  iron. 

The  owners  of  the  property  contracted  with  manufacturers  of  acid  to 
erect  a  plant  to  use  iron  pyrite  from  a  deposit  about  4  miles  from  the 
dumps,  and  to  sell  to  them,  during  a  period  of  10  years,  53°B.  acid  at 
.'$4.32  per  long  ton  (0.19  cents  per  pound).  The  contractors  were  to  ex- 
tract the  copper  from  the  pyrite  paying  the  estate  a  royalty  of  $145  to 
$175  per  ton  of  copper  produced,  and  were  bound  to  burn  4000  to  4800 
tons  of  pyrite  annually.  The  pyrite  was  estimated  to  contain  3.5  to  8.0 
per  cent.  Cu,  and  the  contractors  were  required  to  leach  it  so  as  to  leave 
no  more  than  0.2  to  0.3  per  cent,  copper  in  the  tailings. 

The  method  used  for  the  extraction  of  the  copper  from  the  burned 
pyrite  consists  in  roasting  it  in  a  muffle  with  the  addition  of  sulphuric 
acid  at  a  temperature  of  450  to  550^  C.  This  brings  the  copper  into  soluble 
condition.  The  product  is  then  leached  in  lead-lined  wooden  tanks, 
first  with  water,  then  with  barren  acid  solution  left  after  precipitating 
the  copper  on  iron  plates,  and  finally  with  dilute  acid.  The  copper  is 
precipitated  from  the  soluton  at  boiling  temperature,  on  cast  iron  plates. 
In  this  work  2  lb.  of  acid  and  from  1  to  2  lb.  of  iron  are  used  per  pound  of 
copper  extracted. 

The  oxidized  dump  material  is  composed  of  one-third  large  pieces 
needing  grinding  and  two-thirds  of  fines.  It  is  of  the  following  average 
composition;  SiOj,  37.0  per  cent.;  Fe,  19.6;  Al^Og,  20;  CaO,  0.25;  Cu, 
0.75  per  cent. 

The  ore  is  shoveled  into  side  dump  cars  and  hauled  by  horses  to  the 
end  of  an  inclined  troughed  belt  conveyor  which  raises  it  to  the  top  of 
the  crushing  plant.  Here  it  is  wet  crushed  in  a  breaker  and  Chili  mills 
to  yield  a  pulp  containing  33  per  cent,  dry  solids  and  having  50  per  cent. 



material  fine  enough  to  pass  through  a  136-mcsh  screen.  This  pulp  is 
conducted  by  wooden  launders  to  the  leaching  vats.  Fig.  4.5  represents 
a  general  plan  and  section  of  the  leaching  and  precipitating  plant.  This 
consists  first  of  10  leaching  tanks  184X42X6  1/2  ft.  deep.  These  tanks 
are  erected  on  a  rock  foundation,  and  have  masonry  walls,  covered  with 
4  in.  of  concrete,  then  1  in.  of  reinforced  asphalt.  The  asphalt  covering 
has  lasted  successfully  for  2  years  and  has  been  subjected  to  temperatures 
from  — 40  to  102°  F.  The  plant  is  operated  in  the  warmer  months  only. 
Each  leaching  tank  is  charged  in  8  hours  with  pulp  containing  200 
tons  dry  material.     The  pulp  during  this  time  receives  13.2  tons  of  53°B. 

Elevated  Tallin 

Ckiir  a.pi^rKsoln.  Dleili 

TuillDgfl  Diflcbarge  level  to  Pump 

Section  tkrougli  G-H 

Fig.  45. — Copper  leaching  plant  at  the  Gumeshevesky  Mine,   Russia. 

sulphuric  acid  which  is  run  into  the  tank,  after  which  the  stirring  arms  of 
an  agitator  are  lowered  into  it.  This  machine  is  furnished  with  five 
vertical  shafts,  each  having  a  stirring  arm  at  its  lower  end.  The  agitator 
is  moved  from  end  to  end  of  the  tank  and  stirs  the  pulp  for  9  hours. 
The  material  in  the  tank  is  then  allowed  to  settle,  the  clear  solution  is 
decanted  through  the  launders  to  the  precipitation  tanks.  "Wash  water  is 
run  in  to  fill  the  tank,  the  pulp  is  again  agitated  for  4  hours,  allowed  to 
settle,  and  the  solution  decanted  as  before.  Four  washes  and  decantations 
arc  thus  made.  The  tank  is  provided  with  a  side  door  through  which  the 
tailings  are  discharged,  while  they  are  kept  in  suspension  by  means  of 
the  agitator.  Care  must  be  taken  to  add  sufficient  water  to  make  at 
least  2  1/2  parts  of  water  to  1  of  tailings.  The  tailings  go  to  a  centrifugal 
pump  which  raises  them  to  the  elevated  tailings  discharge.  It  takes 
;5  1/2  hours  to  discharge  a  tank.  The  agitator  is  then  transferred  by  a 
crane  to  another  tank.  The  cycle  of  the  leaching  operation  requires 
5  days,  two  tanks  being  charged  and  two  discharged  daily. 


The  quantity  of  solution  decanted  at  one  time  is  660  tons.  The 
copper  content  of  the  solution  varies  from  0.121  per  cent.  Cu  at  the  first 
decantation  to  0.015  per  cent,  in  the  last  wash.  About  50  per  cent,  of  the 
copper  is  extracted  by  these  leaching  operations,  the  insoluble  balance 
consisting  of  copper  silicate  and  native  copper. 

The  accompanying  table  shows  the  actual  consumption  of  acid  as 
compared  with  laboratory  tests. 


Acid  consumed  by 

Laboratory  tests, 

Actual  practice. 

per  cent. 

per  cent. 






I                  15.0 


I  10.0 



Iron  oxides 


Organic  matter,  etc . 

There  are  20  asphalt  lined  concrete  precipitating  tanks  arranged  in 
five  rows  with  four  tanks  per  row.  Each  tank  is  43  ft.  long,  19  ft.  wide, 
and  2  ft.  7  in.  deep.  The  upper  tank  of  each  row  is  filled  with  110  to  120 
tons  of  cast-iron  plates.  The  other  four  have  inclined  false  bottoms, 
each  covered  with  about  3  tons  of  granulated  cast  iron  in  a  layer  4 
in.  thick.  Thus  there  are  in  all  the  tanks  460  tons  of  plates  and  200  tons 
of  granulated  iron.  The  solution  flows  through  the  system  at  the  rate  of 
nearly  200  tons  per  hour.  The  average  entering  solution  contains  0.04 
to  0.07  per  cent.  Cu,  and  when  leaving  it  contains  0.002  per  cent.,  indicat- 
ing a  precipitation  of  95  per  cent,  of  the  copper.  The  tanks  containing 
the  plates  are  charged  at  intervals  of  3  weeks;  the  other  tanks  every 
4  to  8  days.  The  plates  are  cleaned  by  scraping  and  washing  off  the 
loosely  adherent  copper.  The  cast  iron  granules  are  washed  with  water  in 
a  trommel  perforated  with  1/8-in.  holes.  The  under  size  of  the  trommel, 
consisting  of  cement  copper  and  small  grains  of  iron,  goes  to  a  magnetic 
separator  where  the  iron  is  removed.  The  oversize  of  the  trommel  is 
returned  to  the  tanks.  Granulated  cast  iron  is  in  every  respect  a  more 
convenient  precipitant  than  cast  iron  plates,  bars,  or  scrap  iron. 
Twelve  tons  of  granules  have  the  same  precipitating  capacity  as  120  tons 
of  plates.  It  takes  eight  men  45  hours  to  clean  120  tons  of  plates,  as  com- 
pared with  six  men  taking  care  of  12  tons  of  granules  in  8  hours. 

The  iron  consumed  is  1.8  to  2.0  tons  per  ton  of  copper  recovered. 
The  cement  copper,  containing  60  to  75  per  cent,  of  the  metal,  is  treated 
by  melting  in  a  reverberatory  furnace,  adding  to  it  a  small  quantity  of 
matte  of  the  grade  of  white-metal  to  remove  impurities.     The  resultant 



blister  copper  is  then  rabbled  and  poled  in  the  usual  way  to  produce  a 
brand  of  best-selected  metal. 

The  accompanying  table,  based  on  operations  for  13  days  in  June, 
1909,  gives  an  idea  of  working  costs  under  normal  conditions. 


Item  of  expense 

Wages  and  superintendence 


Assay  office  expense 

Repairs  and  renewals 


Acid,  351  tons  at  $4.32 

Pig  iron,  62.45  tons  at  $10.48 

Refining,  33  tons  at  $12.50 

Tax  on  33  tons  at  $15.96 

Depreciation  and  general  expenses 


Total  cost 

Cost  per  ton  of 

Cost  per  lb.  of 

dump  material 



































From  7735  tons  of  crushed  dump  material,  33  tons,  or  0.43  per  cent, 
of  copper,  was  recovered,  using  341.1  tons  of  53° B.  sulphuric  acid  and 
62.45  tons  of  pig  iron.  The  cost  of  the  plant,  not  including  the  acid 
works,  was  $128,700. 

Ferric  Sulphate,  Ye^iSO^g.- — If  cupriferous  pyrites  is  treated  with  a 
solution  of  ferric  sulphate,  copper  goes  into  solution  in  proportion  to  the 
quantity  of  iron  that  has  been  reduced  from  the  ferric  to  the  ferrous  condi- 
tion. If  ferric  sulphate  is  applied  to  a  fresh  lot  of  cupriferous  pyrites  it  is 
at  first  quite  rapidly  reduced,  and  the  richer  the  mineral  in  copper,  the 
quicker  and  more  perfect  will  be  the  reduction.  As  the  copper  content  in 
the  ore  is  diminished,  the  reducing  action  becomes  slower,  and  in  the  later 
stages  of  treatment,  if  the  ferric  sulphate  solution  is  allowed  to  remain  in 
contact  with  the  partially  exhausted  ore  until  reduction  has  taken  place,  it 
will  be  found,  (1)  that  the  free  sulphuric  acid  is  on  the  increase,  and  (2)  that 
the  copper  content  of  the  solution  does  not  increase  in  proportion  to  the 
quantity  of  ferric  sulphate  reduced,  as  is  the  case  when  it  is  first  applied 
to  untreated  ore.  On  continuing  the  treatment  until  the  reduced  liquor 
shows  no  f urthur  increase  in  its  copper  content,  the  proportion  of  free  acid 
rapidly  increases,  due  to  the  action  of  ferric  sulphate  on  iron  pyrites. 
At  a  slightly  elevated  temperature  the  leaching  action  of  the  mineral  is 
more  rapid. 

Before  the  copper  is  precipitated  with  iron,  precautions  should  be 
taken  to  insure  as  far  as  possible  the  reduction  of  the  ferric  sulphate,  or 
the  consumption  of  iron  will  be  excessive  and  the  precipitation  retarded. 


S.  R.  Adcock  carried  out  at  Rio  Tinto  a  series  of  experiments'  to 
determine  the  amount  of  copper  that  could  readily  be  extracted  from 
cupriferous  pyrites  by  washing  the  mineral  with  a  solution  of  ferric  sul- 
phate, and  also  to  note  the  changes  that  take  place  during  the  opera- 
tion. To  this  end  varying  quantities  of  crude  smalls,  all  of  which  would 
pass  through  a  0.25-in.  sieve,  were  washed  in  small  lead  tanks  with  a 
cold  solution  of  ferric  sulphate  (2  per  cent.  Fe) ;  the  liquor  was  allowed 
to  remain  in  contact  with  the  mineral  until  its  color  indicated  that  the 
greater  part  of  the  iron  had  been  reduced,  when  it  was  drawn  off  and  a 
fresh  quantity  of  ferric  sulphate  added. 

These  experiments  show  how  rapidly  ferric  sulphate  solutions  act 
upon  the  cuprous  sulphide  in  the  pyrites,  and  in  places  where  these 
liquors  are  plentiful  or  can  be  cheaply  manufactured  they  no  doubt 
could  be  used  to  advantage  for  extracting  say  one-half  of  the  total  copper 
content  of  mineral  running  from  1.5  per  cent,  and  upward,  before  it  is 
formed  into  heaps  for  treatment  by  the  open  air  or  weathering  process. 
The  extraction  by  this  process  compares  favorably  with  the  (1)  open-air 
calcination,  until  recently  so  extensively  used  at  Rio  Tinto  and  the  (2) 
"weathering"  or  air  oxidation  process. 

"The  amounts  of  coppei'  extracted  at  different  periods  were  calculated  from 
the  analyses  of  the  liquors,  and  the  results  obtained  from  these  experiments  are 
given  in  the  following  table:  \ 

Weight  of  mineral 

5  kg 

Copper  content 


under  treatment 



Copper  extracted 

4  .  62  per  cent. 

31.90  per  cent. 
44.00  per  cent. 
61.00  per  cent. 
66 .  00  per  cent. 

5  kg 1 .62  per  cent.  4 


2  kg 6.49  per  cent.  I  30 

'  61 

i  :  180 

23.00  per  cent. 
38 .  00  per  cent. 
50 .  50  per  cent. 
60 .  00  per  cent. 

50.70  per  cent. 
63 .  30  per  cent. 
79 .  00  per  cent. 

"The  open-air  calcination  as  applied  to  2.5  per  cent.  Cu  mineral  yields  three- 
fifths  of  the  total  copper  (60  per  cent.)  which  is  at  once  dissolved  out  after  calcina- 
tion.^  The  weathering  method  yields  88  per  cent,  of  the  total  copper  content 
of  the  mineral  treated  in  6  years. ^  While  carrying  out  the  above  trials  it  was 
noticed  in  each  case  that,  after  the  first  two  or  three  washings,  the  liquors  gradu- 
ally increased  in  their  free  acid  content,  and  further  experiments  were  performed 

'Min.  Ind.,  Vol.  IX,  1901. 

2  J.  H.  Collins,  Trans.  Inst.  Min.  and  Met.,  Vol.  II. 

'J.  H.  Brown,  Jour.  Soc.  Chem.  Ind.,  Vol.  XIII. 



to  determine  the  chemical  action  that  was  taking  place.  To  determine  the 
action  of  ferric  sulphate  on  cuprous  sulphide,  20  grm.  of  pure  copper  glance 
(CujS)  crushed  to  a  fine  powder,  was  treated  with  an  excess  of  the  ferric  solution. 
On  filtering  off  the  liquor,  well  washing  the  insoluble  residue,  etc.,  it  was  found 
on  analysis  that  15.39  grm.  of  copper  had  been  dissolved,  27.43  grm.  iron  has 
been  reduced  from  the  ferric  to  the  ferrous  condition,  and  3.83  grm.  of  free 
sulphur  had  been  produced.  Based  on  these  results  it  will  be  seen  that  the  fol- 
lowing equation  represents  the  reaction  that  has  taken  place : 

(1)  Cu  ^S  +  2Fej  (SO J ,  =  2CuS0,  +  4FeS0,  +  S, 

but  haA'ing  noticed  on  every  occasion  when  carrying  out  experiments  on  the 
above  lines,  that  the  first  50  per  cent,  of  the  copper  is  more  easily  extracted,  it?  is 
probable  that  the  reaction  is  more  correctly  shown  as  taking  place  in  two  stages, 

(a)  Cu  ^S  +  Fe,  (SO j',  =  CuS  +  CuSO,  +  2FeS0,. 

(b)  CuS  +  Fe^  (SO.) ,  =  CuSO.  +  2FeS0,  +  S. 

"With  a  view  to  determine  the  action  of  ferric  sulphate  on  iron  pyrites 
FeSj,  100  grm.  low  grade  ore  which  had  been  previously  ground  to  a  fine  powder, 
sampled  and  analyzed,-  was  washed  with  a  strong  solution  of  ferric  sulphate 
(2  per  cent.  Fe)  until  the  analysis  of  the  liquor  drawn  off  from  time  to  time, showed 
that  practically  all  the  copper  had  been  removed.  The  washed  mineral  was  then 
treated  for  63  days  with  the  ferric  solution,  and  at  the  end  of  this  period  it  was 
noticed  that  the  reduction  of  the  ferric  iron  was  taking  place  almost  as  rapidly 
as  at  the  commencement  of  the  experiment.  From  the  analysis  of  the  washings, 
which  were  drawn  off  when  the  color  indicated  that  reduction  had  taken  place, 
it  was  noted  that  each  successive  wash  showed  a  slight  increase  in  its  free  acid 
content  until  the  copper  contained  by  the  mineral  had  been  practically  exhausted; 
at  this  stage  the  free  acid  appeared  to  have  reached  its  maximum,  and  from  thence 
was  always  found  present  in  quantity,  directly  proportional  to  the  amount  of 
ferric  sulphate  that  had  been  reduced  by  the  mineral. 

The  mineral  at  the  conclusion  of  the  experiment  was  well  washed  with  dis- 
tilled water,  and  dried  at  100°  C.  The  following  are  the  analyses  before  and 
after  treatment : 





0.64     per 






51 .  20     per 






43.82     per 






0.56     per 






0.03     per 






0.003  per 






0.97     per 






2.00     per 






0.39     per 






99.613  per  cent. 


per  cent. 

Free  sulphur 


2.51  per  cent. 



"The  above  analyses  show  that  the  copper  and  zinc  originally  contained  by 
the  pyrites  are  almost  totally  extracted  by  ferric  sulphate,  the  arsenic  to  a  lesser 
extent  in  the  same  time,  while  all  the  lead  remains  in  the  washed  material, 
probably  as  insoluble  sulphate.  It  is  well  to  note  that  there  is  free  sulphur  pres- 
ent in  the  material  after  treatment. 

"The  results  of  the  analyses  of  the  liquors  obtained  by  washing  the  mineral 
for  63  days,  after  the  copper  had  been  extracted,  were  as  follows:  During  this 
trial  the  liquors  were  kept  at  a  slightly  elevated  temperature;  45.2  grm.  of  ferric 
iron  reduced  to  the  ferrous  state;  44  grm.  free  sulphuric  acid  formed,  and  4.2 
grm.  of  iron  (from  the  pyrites)  dissolved.  Based  on  these  results  it  is  probable 
that  the  following  equation  represents  the  action  of  ferric  sulphate  on  iron  pyrites : 

(2)   1  IFe^  (SO  J  3  +  2FeS3  +  UTIfi  =  24FeSO,  +  I2H2SO,  +  S. 

"  The  amount  of  pyrites  required  to  effect  the  reduction  is  very  small,  but  the 
reaction  in  the  cold  is  slow  in  taking  place,  unless  a  large  excess  of  mineral  is 
exposed  in  proportion  to  the  quantity  of  ferric  sulphate  to  be  reduced;  at  a 
temperature  of  from  50  to  60°  C.  the  reaction  takes  place  much  quicker  than  at 
ordinary  temperatures." 

Ferric  sulphate  also  reacts  with  cupric  oxide,  according  to  the  equation : 

(3)  3CuO+Fe2(SOj3  =  3CuSO,+Fe203, 

and  somewhat  similarly  on  the  carbonate,  Vfith  the  liberation  of  CO2. 
Zinc  in  its  sulphide  combination  is  readily  soluble  in  a  solution  of 
ferric  sulphate,   according  to  the  equation: 

(4)  ZnS-|-Fe2(S04)3  =  ZnSO,  +  2FeS04  +  S. 

From  equations  1,  3,  and  4,  it  is  readily  estimated  that  theoretically, 
6.3  lb.  of  anhydrous  ferric  sulphate  are  required  to  dissolve  1  lb.  of  copper 
from  its  cuprous  sulphide  combination,  and  2.1  lb.  from  the  cupric  oxide: 
6.1  lb.  of  ferric  sulphate  is  destroyed  in  dissolving  1  lb.  of  zinc  from 
sphalerite,  and  as  zinc  is  more  electropositive  than  iron,  it  will  not  be 
precipitated  by  it,  but  will  accumulate  indefinitely  unless  some  means  is 
provided  to  remove  the  zinc  sulphate,  or  at  intervals  renew  the  solution. 

Experiments  with  Ferric  Sulphate  at  Cananea.— W.  L.  Austin  gives  an 
account^  of  experiments  carried  out  at  Cananea,  State  of  Sonora,  Mexico, 
by  the  Cananea  Consolidated  Copper  Co. ,  from  the  original  notes  furn- 
ished by  David  Cole,  assistant  general  manager  of  the  company,  with  a 
view  to  ascertaining  the  leaching  qualities  of  local  material.  While  these 
experiments  did  not  lead  to  the  adoption  at  Cananea  of  the  method  ad- 
vocated, the  results  obtained  are  instructive  and  of  practical  value  to 
any  one  contemplating  similar  experimental  work. 

"The  material  treated  at  Cananea  consisted  of  mill-tailing  sands  and  of  flue 
dust  from  the  furnaces.  The  leaching  was  done  by  simple  percolation  without 
agitation.     The  copper  was  precipitated  from  the  cuprous  liquors  by  means  of 

'  "Mines  and  Methods,"  Sept.,  1910. 


metallic  iron.  The  spent  solutions  were  regenerated  by  forcing  heated  air 
through  them.  The  principal  difficulties  encountered  arose  in  the  regeneration 
of  the  liquors. 

"There  is  from  0.27  to  0.66  per  cent,  zinc  in  the  mill-tailing  sands  at  Cananea 
and  the  copper  varies  from  0.54  to  0.89  per  cent.  It  was  found  that  the  zinc 
minerals  were  more  readily  attacked  than  those  carrying  the  copper,  for  in  the 
same  time  that  approximately  65  per  cent,  of  the  copper  was  extracted  72  per 
cent,  of  the  zinc  went  into  solution.  It  was  thought  at  Cananea  that  dilution 
of  the  solutions  by  wash  water  introduced  to  remove  the  soluble  copper  salts 
from  the  tails  before  allowing  them  to  go  to  waste  would  suffice  to  keep  the  zinc 
content  in  the  solution  low  enough  to  prevent  serious  fouling.  A  certain  propor- 
tion of  the  zinc  salts  would  be  removed  in  this  manner  because  the  wash  water 
after  flowing  over  iron  to  deposit  the  copper  would  be  allowed  to  escape,  but  the 
effect  of  zinc  salts  in  the  apphcation  of  the  process  in  any  particular  case  can  only 
be  determined  by  experimenting  with  the  ore  in  question  on  a  considerable  scale. 
In  using  2  lb.  solution  to  1  lb.  of  ore  it  was  found  that  94  per  cent,  of  the  soluble 
matter  is  extracted  without  increasing  the  bulk  of  solution  with  wash  water. 

"  Only  very  small  amounts  of  alkalis  went  into  solution  in  treating  the  sand 
tailings — about  0.1  per  cent. — and  the  quantities  of  alumina  and  lime  were 
negligible.  The  pyrites  in  the  ore  was  found  to  reduce  the  ferric  salts,  cau.sing  a 
waste  of  the  solvent. 

"It  was  found  that  the  average  loss  of  ferric  sulphate  per  pound  of  copper 
extracted  was  4.37  lb.  This  figure  was  derived  by  crediting  the  gain  in  ferrous 
sulphate  from  both  the  action  of  the  ore  and  from  the  precipitation  of  the  iron, 
and  'calculating  the  oxidation  in  the  tower.' 

"It  was  found  at  Cananea  that  a  ferric  sulphate  solution  accomplished  a  very 
complete  extraction  of  the  copper  from  material  containing  the  oxides  and 
carbonates,  but  that  it  acted  more  slowly  upon  those  carrying  the  metal  in  the 
form  of  chalcocite.  Chalcppyrite  was  hardly  attacked  at  all.  The  content  of 
the  solution  in  ferric  sulphate  was  found  to  be  of  comparatively  small  importance 
provided  base  salts  were  absent.  It  worked  when  it  carried  as  small  an  amount 
of  ferric  salt  as  1  per  cent.,  and  the  extractions  were  nearly  as  complete  when 
using  2  per  cent,  solution  of  ferric  sulphate  as  when  using  7  per  cent.  A  freshly 
precipitated  solution  acted  more  energetically  than  an  old  one,  even  though 
relatively  weaker.  The  presence  of  much  basic  sulphate  was  found  to  greatly 
retard  the  leaching,  'clogging'  the  action.  For  this  reason  the  solution  was 
settled  in  the  oxidizer  before  using.  About  65  per  cent,  of  the  copper  contents 
of  the  mill-tailing  sands  could  be  extracted  in  about  three  hours  when  the  liquor 
was  boiled,  or  in  about  7  hours  when  it  was  kept  at  a  temperature  of  70° 

"Concentrates  gave  about  40  per  cent,  extraction. 

"After  the  liquors  were  thought  to  contain  a  sufficient  quantity  of  copper  in 
the  sulphate  form  they  were  conveyed  to  the  precipitating  tanks  where  the 
metal  was  removed  by  being  brought  in  contact  with  metallic  iron : 

(5)  CuSO,+Fe=FeSO,  +  Cu. 

"The  consumption  was  light  because  when  the  solution  is  hot  the  precipita- 
tion is  very  rapid  and  complete. 

"After  the  liquors  had  passed  through  the  precipitating  boxes  the  iron  was 


practically  all  in  the  ferrous  condition  and  it  became  necessary  to  regenerate  the 
solvent — that  is,  to  reoxidize  the  iron — before  it  would  again  be  available.  It 
is  precisely  in  the  regeneration  of  the  spent  solutions  that  the  weakness  of  leach- 
ing methods  based  on  the  use  of  ferric  salts  becomes  apparent. 

"  At  Cananea  the  regeneration  was  effected  by  pumping  the  solution  into  an 
oxidizing  tank  where  it  was  heated  by  a  steam  coil  and  agitated  and  oxidized  by 
hot  air.  The  capacity  of  the  oxidizer  was  30,000  lb.  of  solution;  the  working 
height  of  the  column  was  11  ft.  8  in.  It  was  thought  that  18  ft.  would  give  better 
results.  From  the  oxidizer  the  solution  was  led  to  a  settler  and  from  which  the 
clear  liquor  was  drawn  off  to  a  feed  tank.     It  was  then  available  for  leaching. 

"The  reactions  which  take  place  when  an  attempt  is  made  to  oxidize  ferrous 
sulphate  to  the  ferric  condition,  without  the  presence  of  free  acid,  are  very 
complicated.  Basic  ferric  salts,  of  which  there  are  many  varieties,  invariably 
form  and  are  precipitated,  thereby  causing  the  loss  of  a  large  part  of  the  iron 
unless  free  sulphuric  acid  has  been  added  in  the  amounts  necessary  to  produce 
the  neutral  ferric  sulphate.  For  the  purpose  specified  (the  formation  of  neutral 
ferric  sulphate),  ten  parts  of  ferrous  sulphate  require  two  parts  of  concentrated 
sulphuric  acid.  The  reactions  which  occur  in  the  transformation  are  indicated 
by  the  following  equation: 

(6)  2(FeSOJ+H2SO,  +  0=Fe2(SOj3  +  H20. 

"If  a  solution  of  neutral  ferric  sulphate  is  heated  with  ferric  hydrate,  there 
results  a  deep  brown  liquor  containing  a  more  basic  salt — two-thirds  as  much 
sulphuric  acid  combined  with  the  same  amount  of  iron  as  in  the  neutral  salt. 
This  basic  salt  is  also  formed  when  a  solution  of  ferrous  sulphate  is  slowly  oxi- 
dized by  contact  with  the  air,  while  at  the  same  time  a  still  more  basic  salt  is  pro- 
duced (with  one-sixth  as  much  sulphuric  acid  as  is  present  in  the  neutral  sulphate) , 
together  with  other  soluble  sulphates  and  the  neutral  ferric  sulphate. 

"  The  basic  salt  containing  two-thirds  as  much  sulphuric  acid  as  is  present  in 
the  neutral  sulphate,  is  decomposed  by  heating,  or  by  dilution  of  the  solution,  the 
resulting  products  being  neutral  ferric  sulphate  and  a  yellow  precipitate  con- 
taining the  one-sixth  salt  referred  to  above.  These  two  last  named  ferric  com- 
pounds predominate  when  a  solution  of  ferrous  sulphate  is  oxidized  by  exposure 
to  the  air,  and  are  claimed  by  some  authorities  to  be  the  final  products  from  the 
oxidation  described. 

.(7)   10(FeSOJ  +  O,=3(Fe,(SO,)3)+Fe,SO,. 

"From  the  above  equation  it  is  evident  that  in  converting  a  ferrous  into  a 
soluble  ferric  sulphate  by  oxidation  in  the  air,  with  the  assistance  of  heat,  without 
the  addition  of  free  acid,  40  per  cent,  of  the  iron  is  deposited  in  the  form  of  insolu- 
ble basic  ferric  sulphates  and  is  therefore  lost  as  an  active  reagent  for  the  required 
purposes,  in  addition  to  a  loss  of  the  acid  with  which  it  is  combined.  Hence  in 
order  to  keep  up  the  grade  of  the  solvent  in  leaching  with  ferric  salt,  it  becomes 
necessary  to  constantly  replenish  the  acid  constituent  of  the  compound,  the 
latter  being  the  reagent  consumed,  and  the  process  becomes  virtually  one  of 
leaching  with  sulphuric  acid.  Where  a  heavy  iron  sulphide  is  calcined  before 
leaching,  the  loss  of  iron  becomes  of  comparatively  small  importance,  but  where 
the  ferrous  salt  is  an  item  of  expense,  the  cost  is  considerable.     If  10  lb.  of  ferrous 

( 'HEMIC.  1 L  PROi 'EsSSES 


sulphate  are  required  to  carry  out  a  certain  leaching  operation,  4  lb.  will  be  lost 
in  the  process  of  regeneration  by  aeration,  unless  2  lb.  of  concentrated  (66°  B.) 
sulphuric  acid  is  added  to  the  solution  to  prevent  the  iron  from  taking  the  form 
of  a  basic  salt.  Therefore,  in  estimating  the  cost  of  leaching  a  given  ore,  the 
relative  expense  in  providing  free  acid,  as  against  that  for  ferrous  sulphate,  has 
to  be  considered. 

"From  the  above  remarks,  it  is  apparent  that  the  reactions  which  take  place 
in  the  regeneration  of  the  spent  solutions  by  the  methods  used  at  Cananea  are  too 
complicated  to  be  written  down  in  formulae;  the  actual  figures  disclosing  results 
obtained  are  more  illuminating. 

"In  making  the  ferric  solution  employed  at  Cananea  the  ferrous  sulphate 
used  contained  sulphide  impurities  which  caused  some  irregularities  during  the 
oxidation,  the  ferric  sulphate  being  destroyed.  The  sulphides  might  have  been 
removed  by  dissolving  the  sulphate  in  a  separate  tank  and  decanting  the  clear 
solution.  The  ferrous  salt  was  added  to  hot  water  into  which  had  been  pumped  , 
some  ferrous  solution  and  the  whole  was  heated  by  means  of  a  steam  coil  near  the 
bottom  of  the  oxidizing  tank.  Air  was  forced  in  by  a  compressor  after  passing 
through  a  heater  where  is  attained  a  temperature  of  from  200  to  400°  F.  and 
served  to  agitate  and  oxidize  the  solution. 

"After  it  was  apparent  that  the  ferrous  salt  had  gone  into  solution,  samples 
of  the  liquor  were  taken  which  gave  the  following  results : 




per  cent. 

per  cent. 




degrees  C. 

















































Total  ferric  sulphate  formed  in  32  hours  was  2100- 
formed  in  1  hour  =  50.6  lb. 

-480=1620  lb.     Average  ferric  sulphate 

"The  table  shows  no  oxidation  in  the  solution  between  the  fifth  and  sixth 
samples.  A  possible  explanation  of  this  feature  is  that  ferric  sulphate  attacked 
the  sulphide  of  iron  already  referred  to  as  an  impurity  in  the  ferrous  salt  fed  to 
the  oxidizing  tank. 

"A  basic  sulphate  which  formed  in  the  tank  during  the  process  of  oxidation 
is  said  to  have  interfered  to  a  great  extent  with  the  operation. 

"In  the  test  from  which  the  above  results  were  obtained,  300  lb.  sulphuric 
acid  were  added  after  the  solution  had  been  oxidizing  for  28  hours.  Four  hours 
later  a  valve  leaked  and  the  solution  in  the  tank  had  to  be  drawn  off,  so  that  it 
was  thought  that  the  basic  sulphate  may  not  have  received  the  full  benefit  of 



the  acid.  The  iron  sulphides  in  the  solution  were  considered  to  have  retarded 
oxidation  to  a  marked  extent. 

"At  Cananea  the  oxidation  of  the  ferrous  to  the  ferric  sulphate  was  found  to 
be  a  serious  problem  as  the  transformation  was  'i'ery  slow,  and  it  was  thought 
that  a  successful  application  of  the  process  will  depend  more  on  a  satisfactory- 
solution  of  this  feature  than  on  any  other. 

"In  an  experiment  made  previous  to  the  one  mentioned,  before  alterations 
had  been  made  in  the  oxidizing  tank,  6.4  tons- of  solution  were  treated  for  2  days. 
During  that  interval  1,500,000  cu.  ft.  of  air  at  a  temperature  of  about  140°  C. 
(equivalent  to  about  100,000  lb.)  were  forced  through  the  solution  with  the  follow- 
ing results; 

Before  blowing. 
After  blowing. . 

Ferrous  sulphate 

Ferric  sulphate 

Per  cent. 



Per  cent.        i         Pounds 


"It  is  stated  that  only  125.44  lb.  of  ferric  sulphate  were  formed,  which  also 
included  the  precipitated  basic  salts.  If  all  the  oxygen  in  the  air  had  been  util- 
ized, nearly  200,000  lb.  ferric  sulphate  might  have  been  produced  had  the 
necessary  amount  of  ferrous  solution  been  available.  It  took  1200  times  more 
air  than  was  theoretically  necessary  to  produce  the  results  desired. 

"In  order  to  increase  the  efficiency  of  the  oxidizer,  perforated  discs  were 
submerged  in  the  solution  and  heated  air  from  the  compressor  was  forced  in  at 
the  bottom.  The  improved  oxidizer  gave  much  better  results  than  were  obtained 
with  the  former  one. 

In  making  an  estimate  of  the  expense  of  converting  the  ferrous  solution  at . 
Cananea,  no  account  appears  to  have  been  kept  of  the  ferrous  salt,  nor  of  labor 
and  repairs;  only  the  following  items  are  given: 

Steam  for  heating  the  solution  through  lead  coil. 
Power  used  for  compressor. 
Coal  to  heat  the  air,  262  lb. 
Sulphuric  acid,  300  lb.. 




"These  figures  are  for  24  hours'  run.  The  sulphuric  acid  is  taken  at  $0.0066 
per  pound,  at  which  price  it  was  thought  that  probably  it  could  be  manufactured 
at  Cananea. 

"As  the  average  amount  of  ferric  sulphate  formed  in  the  oxidizer  was  50.6  lb. 
per  hour,  dividing  the  17.70  by  24  to  obtain  the  cost  per  hour,  and  dividing  this 
result  by  50.6  to  get  the  cost  per  pound,  gives  $0.0063  as  the  cost  of  1  lb.  of  ferric 
sulphate  produced  in  the  manner  described. 

"An  estimate  made  by  the  engineers  who  conducted  the  tests  at  Cananea  is 
given  below.     These  figures  are  said  to  represent  the  actual  results  obtained  in 

Per  ton 

Per  cent. 






2.5 . 5 












treating  a  lO-ton  lot  of  Cobre  Grande  ore.  This  ore  contained  3  percent,  copper, 
and  the  extraction  is  said  to  have  been  96  per  cent. 

Crushing  to  six  mesh, 



Acid,  16  lb.  at  1/2  cent. 

Coal  to  evaporate  wash  water. 

Iron  to  precipitate  copper. 

Heating  solution. 


Total,  $3.73  100.00 

ex  "This  is  equivalent  to  $0,064  per  pound  of  copper.     It  was  found  in  this 
perimental  work  that  the  best  leaching  results  were  obtained  when  the  copper 
was  in  the  oxidized  condition.     A  test  made  on  1735  lb.  of  flue  dust  containing 
5.63  per  cent,  copper  was  satisfactory,  as  shown  below. 

"The  dust  was  first  calcined  to  oxidize  the  suljjhides  and  then  leached  with 
a  hot  solution  of  ferric  sulphate.  The  leaching  proceeded  rapidly;  two  applica- 
tions of  the  solution  left  practically  no  copper  in  the  tailings.  The  following 
table  shows  the  consumption  of  ferric  sulphate  in  this  operation: 


Ferric  sulphate  (Fe2(S04)3)  in  feed  solution,  440 

Ferric  sulphate  (Fe2(SO.,)3)  in  tail  solution,  190 

Ferric  sulphate  (FejCSOj),)  used  to  leach,  250 

Copper  in  charge,  97 . 5 

Copper  n  solution,  95  per  cent,  extraction,  92 . 6 

Ferric  sulphate  consumed  per  pound  of  copper  extracted,  2 . 7 

The  time  factor  was  ignored  in  this  test,  the  perfection  of  the  leach  alone  being 
considered.  The  dust  contained  about  30  per  cent,  iron  so  that  an  excess  of 
ferrous  sulphate  was  found  in  the  tail  solution.  No  attempt  was  made  at  esti- 
mating the  cost  in  these  tests  because  the  expense  for  calcining,  labor,  steam  to 
heat  the  solutions,  repairs,  etc.,  could  not  be  accurately  obtained. 

"Another  test  said  to  have  been  made  on  a  10-ton  lot  of  flue  dust  assaying 
7.5  per  cent,  copper,  resulted  in  an  extraction  of  94  to  96  per  cent.  The  cost 
items  were  given  as  follows ; 

Per  ton 

Per  cent. 







Acid,  16  lb.  at  1/2  cent. 



Iron  for  precipitating  the  copper. 



Steam  for  heating  solufon 



Coal  for  evaporating  the  wash  water, 






Total,  $4 .  17         100 . 0 

Cost  per  pound  of  copper,  $0,029. 



"  The  Cananea  tests  appear  to  have  come  to  a  sudden  termination  through  the 
breaking  down  of  the  oxidizer,  without  the  cost  of  the  solvent  having  been  con- 
clusively established.     The  extraction  was  apparently  satisfactory. 

"The  figures  given  in  the  foregoing,  relative  to  the  cost  of  producing  the  ferric 
sulphate  are  either  based  upon  what  might  have  been  accomplished  with  a  per- 
fect oxidizer,  or  are  incomplete,  or  are  theoretical  deductions  from  imperfect 
results.     They  are  not  conclusive. 

"  It  is  to  be  regretted  that  having  carried  the  work  along  to  the  point  reached, 
an  effort  was  not  made  to  definitely  ascertain  the  cost  of  producing  the  solvent, 
or  what  is  the  same  thing,  the  expense  of  regenerating  the  solutions. 

"An  estimate  made  by  an  engineer  in  the  early  stages  of  the  experiments  as 
to  the  probable  cost  of  producing  copper  by  this  method  is  given  below.  This 
estimate  was  figured  before  improvements  in  the  oxidizer  were  made.  No 
charges  are  included  for  mining,  transportation,  crushing,  calcining,  general 
expenses,  etc. 



Per  ton  sands 

Per  pound 

Per  cent. 




Power,  at  $0.65  per  horse-power  day  for  blowing  air 

Acid — if  made  on  ground — 2  lb.  acid  per  1  lb.  Cu 



Shipping  and  refining  at  $0.05  per  pound  copper 

Added  for  waste  of  material,  20  per  cent,  of  $0.89 






"A  summary  of  the  cost  of  producing  the  ferric  sulphate  used  in  the  experi- 
ments at  Cananea,  derived  from  independent  sources,  is  given  below.  The 
figures  include  only  steam  heat,  heat  for  air,  power  for  compressor,  and  acid. 


Lb.  ferric  sulphate  consumed 

Ferric  sulphate  cost 

Ferric  sulphate  cost  per  pound 

per  pound  of  copper 

of  1  lb. 

of  copper 



















"The  cost  per  pound  of  ferric  sulphate  in  No.  II  is  probably  the  closest 
approximation  to  the  actual  cost  of  the  pound  of  ferric  sulphate  made  at 
Cananea,  but  does  not  include  original  cost  of  the  ferrous  sulphate,  losses  in 
handling,  basic  salts,  nor  losses  from  presence  of  impurities." 


Thomas'   Experiments  with   Ferric   Sulphate  on  Sulphide  Ore.' — F. 

Thomas  experimented  with  ore  of  the  following  composition: 

SiOj,  gangue,  44 . 7     per  cent. 

Cu^S  (10.5  per  cent.  Cu),  13.5    per  cent. 

Fe^Sj,  25.25  percent. 

FejOa,  5 .  76  per  cent. 

MnO,  0.18  percent. 

AljOj,  7.5     percent. 

PzOs,  0 .  34  per  cent. 

CaO,  2 .  83  per  cent. 

Traces,  Sb,   Sr,  Mg,  and  K. 

The  principal  results  of  his  experiments  are  as  follows:  The  double 
sulphides  of  copper,  which  occur  in  nature,  require  for  their  complete 
transformation  by  means  of  ferric  sulphate  such  a  long  treatment  and 
so  fine  crushing,  that  a  commercial  application  of  this  method  of  leaching 
does  not  pay  under  the  conditions  existing  in  most  copper  producing 
countries.  Copper-iron  sulphides,  artificially  prepared,  also  resist  the 
action  of  ferric  sulphate  in  the  same  way  as  the  natural  ones.  Free 
copper  sulphides  and  oxides  react  with  ferric  sulphate  in  aqueous  solution 
easily  and  quickly.  The  reason  to  which  the  difficulty  of  the  action  of 
ferric  sulphate  in  the  former  cases  is  due  must,  therefore,  lie  in  the  chem- 
ical affinity  between  the  ingredients  of  the  natural  and  artificial  sulphur 
ores  in  copper.  The  presence  of  a  larger  quantity  of  ferrous  sulphate  in 
the  solution  impairs  the  solution  of  copper  from  cuprous  sulphide  by 
means  of  ferric  sulphate.  The  method  suggested  in  the  Siemens-Halske 
process  for  roasting  copper  ore,  so  that  the  main  quantity  of  the  iron  is 
transformed  into  oxide  while  the  main  quantity  of  copper  remains  as 
cuprous  sulphide,  is  practically  impossible.  Neither  can  satisfactory 
results  be  obtained  by  means  of  dead-roasting  of  sulphide  ores,  for  the 
reason  that  at  the  temperature  required  for  this  purpose  basic  silicates 
are  formed  by  means  of  combination  of  the  copper  oxide  with  the 
silicates  of  the  gangue,  and  that  perhaps  also  salts  of  theFe304type  are 
formed  by  the  combination  of  the  oxides  of  copper  and  iron;  such  salts 
are  acted  upon  very  slowly  by  ferric  sulphate.  The  double  sulphides 
must  therefore  be  destroyed  by  oxidizing  roasting  at  so  low  a  temperature 
that  the  formation  of  the  compounds  just  mentioned  is  impossible.  This 
temperature  is  about  450  to  480°  C.  The  product  thus  obtained  contains 
the  copper  essentially  in  the  form  of  sulphate.  The  low  temperature  of 
roasting  permits  the  use  of  simple  furnaces,  and  a  crushing  of  the  ore 
is  sufficient,  corresponding  to  484  meshes  per  square  centimeter.  The 
leaching  can  be  so  conducted  that  a  solution  of  copper  sulphate  is  obtained 
with  only  small  quantities  of  iron. 

'  Metallurgie,  Jan.  15,  Feb.  8,  and  22,  1904. 


Experiments  in  Southern  Tyrol,  Spain.' — A  copper  plant  was  erected 
in  Southern  Tyrol,  in  which  an  attempt  had  been  made  to  employ  the 
old  Siemens-Halske  process,  the  first  stage  of  which  is  leaching  with 
ferric  sulphate  solution.  This  preliminary  leaching  was  a  failure,  so 
that  the  subsequent  electrolytic  precipitation  of  the  copper  was  never 
attempted.  The  reason  was  that  the  ore  contained  the  copper  in  the 
form  of  the  compound  Cu2S,FeS,FeS2;  this  was  roasted  at  such  a  high 
temperature  that  the  copper  oxide  combined  with  other  oxides  and 
formed  combinations  like  CuO,Fe203  and  silicates,  which  are  not  amen- 
able to  leaching.  The  ore  was  afterward  roasted  at  a  low  temperature, 
which  was  then  easily  leached,  and  it  was  decided  to  work  the  copper 
ore  simply  for  copper  sulphate. 

Copper  Extraction  at  Kedabeg,  Russia.^ — The  rich  ores  at  Kedabeg 
are  smelted.  The  lean  ores,  containing  less  than  5  per  cent,  copper, 
say  3  per  cent.,  and  which  consequently  would  not  bear  the  cost  of 
direct  smelting,  are  treated  by  leaching.  The  process  is  very  simple 
and  well  adapted  to  the  local  conditions  which  scarcely  permit  of  the 
use  of  salt  or  other  decomposing  reagents.  The  ore  is  cheaply  roasted 
in  kilns  or  Gerstenhofer  furnaces  without  fuel,  the  copper  being  thus 
brought  into  soluble  form  for  leaching.  This  presents  no  great  diffi- 
culty, as  the  copper,  originally  existing  as  sulphide,  is  oxidized  par- 
tially to  sulphate  in  the  furnace,  the  sulphatization  being  completed 
to  a  certain  extent  later  during  the  leaching  by  the  ferric  sulphate 
formed  in  the  roasting  and  also  present  in  the  residual  solutions 
from  the  electrolytic  precipitation,  the  latter  being  run  into  leaching 
ponds.  A  comparatively  long  time  is  required  to  obtain  a  practically 
complete  change  of  the  copper  to  sulphate,  several  years'  leaching  being 
necessary  to  reduce  the  copper  contents  to  0.5  or  0.7  per  cent,  copper. 
However,  from  50  to  70  per  cent,  is  obtained  at  a  very  small  cost  in  the 
first  year.  The  whole  plant  is  in  the  open,  without  covering  or  roofing, 
and  on  sloping  ground.  The  ground,  very  impermeable  to  begin  with,  is 
completely  hardened  by  the  decomposition  of  basic  salts  of  iron. 

The  copper  is  precipitated  in  wooden  tanks  by  means  of  scrap  iron,  a 
rapid  circulation  being  kept  up  all  the  time.  In  this  way  409.5  metric 
tons  of  cement  copper,  with  65  to  75  per  cent.  Cu,  are  produced  per 

The  Millberg  Process.^ — Well  roasted  copper  pyrites  or  copper  cinders 
contain  copper  as: 

1.  Copper  sulphate,  soluble  in  water. 

2.  Copper  oxide,  soluble  in  ferric  sulphate. 

'  W.  Borohers,  Metallurgie,  Aug.  8,  1909. 
"  Gustave  KoUe,  Min.  Ind.,  Vol.  VI. 
'  Chemicker  Zeitung,  XXX,  511. 


3.  Cuprous  oxide,  capable  of  being  oxidized  by  ferric  sulphate  which 
will  then  dissolve  it. 

4.  Copper  sulphide,  oxidizable  by  ferric  sulphate  which  will  then 
dissolve  it. 

Millberg's  method  consists  in  leaching  the  roasted  ore  or  cinders  with 
a  solution  of  ferric  sulphate  by  which  the  copper  is  salts  are  dissolved 
out  and  pass  into  the  filtrate.  This  filtrate  will  then  contain  ferric 
sulphate,  ferrous  sulphate,  copper  sulphate  and  sulphate  of  other  metals 
when  present,  such  as  zinc,  manganese,  cobalt,  nickel,  and  aluminum. 
Ferric  sulphate  is  very  effective  in  bringing  the  copper  salts  into  solution 
so  that  in  burnt  pyrites  containing  0.8  per  cent,  to  4.0  per  cent,  copper, 
there  is  left  in  the  residues  0.05  per  cent,  to  0.2  per  cent.  onlj'. 

The  filtrate  contained  in  a  tank  is  brought  to  the  temperature  of  60°  C. 
and  to  it  is  added  a  little  ferric  sulphate  solution  till  by  testing  with 
ferrocyanide  solution  the  end  point  is  reached.  This  may  be,  for  example, 
0.66  per  cent,  of  a  ferric  sulphate  solution  of  25°  B.  This  oxidizing 
action  will  not  exceed  2  days,  and  for  some  kind  of  cinders  but  a  few 
hours.  Air  is  then  blown  into  the  solution  in  presence  of  an  alkali  base, 
such  as  milk  of  lime.  The  ferric  sulphate  is  precipitated  as  an  insoluble 
basic  sulphate.  This  is  filtered  or  decanted  off  and  there  remains  in  the 
filtrate  the  sulphates  of  copper  and  other  metals. 

The  solution  is  heated  to  the  boiling  point,  and  to  it  is  slowly  added 
dilute  milk  of  lime,  which  precipitates  the  copper  as  an  insoluble  basic 
sulphate  of  a  light  green  color  and  leaves  the  other  sulphates  in  solution. 
The  precipitation  must  be  watched  and  the  treatment  stopped  as  soon  as 
the  copper  has  been  precipitated. 

The  Elliott  Process.' — The  Elliott  process  consists,  essentially,  in 
leaching  the  ore  with  a  hot  non-acid  solution  of  ferrous  sulphate,  passing 
air  through  the  solution  during  the  operation  of  leaching,  precipitating 
the  copper  from  the  solution  with  iron,  thereby  regenerating  the  ferrous 
sulphate  for  a  repetition  of  the  process. 

The  process  is  applicable  to  oxidized  ores;  if  the  ore  to  be  treated  is  a 
sulphide,  it  has  to  be  given  a  preliminary  roasting  to  convert  the  sulphide 
into  the  oxide  or  sulphate.  The  air,  preferably  heated,  converts  the 
ferrous  sulphate  to  the  ferric  sulphate,  which  then  acts  on  the  copper 
oxide.  The  oxidation  of  the  ferrous  sulphate  to  ferric  sulphate  may  be 

6FeS04  +  03  =2Fe2(SOj3-|-re203 

the  leaching  operation: 

3CuO  +Fe2(S0  J  ^  =  3CuS0,  +Fe,03 

and  the  precipitation: 

CuS04+Fe  =  Cu+FeS0, 
'U.  S.  Patent  814,836,  March,  1906. 


so  that  the  leaching  solution,  is  therefore  regenerated  by  the  precipitation 
of  the  copper  with  iron.  The  ferrous  sulphate  solution  is  then  again 
applied  to  the  ore,  and  while  hot,  air  is  passed  through  it,  thereby  re- 
generating the  ferric  sulphate,  which  attacks  the  copper  and  makes  it 
soluble,  thus  repeating  the  cycle.  It  is  claimed  that  the  process  is 
applicable  to   ores   containing  too   much  lime  for  acid  treatment. 

The  Laist  Process. — This  process  is  based  on  the  use  of  sulphuric  acid 
as  the  solvent,  and  hydrogen  sulphide  as  the  precipitant. 

The  steps  in  the  process  may  be  summarized  as  follows : 

1.  Dissolving   copper  from  the   ore  with   dilute   sulphuric   acid. 

2.  Precipitating  the  copper  from  its  solution  as  copper  sulphide  by 
hydrogen  sulphide,  accompanied  by  the  regeneration  of  acid. 

3.  Manufacture  of  the  hydrogen  sulphide  gas:  a.  Reduction  of 
gypsum  to  calcium  sulphide  with  coal;  b.  Decomposition  of  the  calcium 
sulphide  by  carbon  dioxide  and  water  to  hydrogen  sulphide  and  calcium 

4.  Conversion  of  the  copper  precipitate  to  metallic  copper. 
The  process  is  based'  on  the  following  reactions: 

(1)  CuC03  +  H2S04  =  CuS04-|-C02  +  H20. 

(2)  CuS04  +  H2S  =  H2S04-l-CuS. 

(3)  CaS04  +  4C  =  CaS+4CO. 

(4)  CaS+H20  +  C02=CaC03  +  H2S. 

The  first  well-known  reaction  shows  what  takes  place  when  copper 
carbonate  or  oxide  is  dissolved  with  sulphuric  acid.  Reaction  2  shows 
the  effect  of  the  hydrogen  sulphide  on  the  copper  sulphate,  by  which  the  cop- 
per is  precipitated  as  sulphide  and  an  equivalent  amount  of  sulphuric  acid 
regenerated.  The  third  reaction,  is  the  first  step  in  making  the  hydrogen 
sulphide  gas.  Calcium  sulphate  (gypsum)  is  reduced  with  coal.  This 
reaction  takes  place  at  a  bright  red  heat,  or  about  1800°  F.  The  reaction 
is  accompanied  by  the  formation  of  both  carbon  dioxide  and  carbon 
monoxide,  but  the  monoxide  largely  predominates.  In  the  reaction  it  is 
assumed  that  carbon  monoxide  only  is  formed.  From  the  fourth  reaction 
it  is  seen  that  when  calcium  sulphide  is  treated  Avith  water  and  carbon 
dioxide,  it  decomposes,  and  calcium  carbonate  and  hydrogen  sulphide  are 

From  the  reactions  it  is  clear  that  168  parts  of  gysum  will  furnish 
enough  hydrogen  sulphide  to  precipitate  63  parts  of  copper.  For  the  reduc- 
tion of  this  gypsum  48  parts  of  carbon  is  required.  In  practice  there  is  used 
3  parts  of  coal  to  7  parts  of  anhydrous  calcium  sulphate,  that  is  to  about 
8  parts  of  gypsum.  This  slight  excess  of  coal  is  necessary  to  effect  a 
quantitative  reduction.  Hence  about  4  lb.  of  the  mixture  of  coal  and 
gypsum  are  used  in  precipitating  1  lb.  of  copper,  or  about  2  2/3  lb.  of 
gypsum,  or  1  1/4  lb.  of  calcium  sulphide  precipitate  1  lb.  of  copper. 


The  copper  sulphide  precipitate  is  readily  converted  into  metallic 
copper,  while  sulphur  dioxide  is  released. 

Most  copper  ores  contain  substances  besides  copper  which  consume 
acid.  Acid  lost  in  this  way  is  not  recovered  in  the  precipitating  tanks, 
but  is  recovered  by  allowing  the  acid  solution  itself  to  flow  down  con- 
densers up  which  sulphur  dioxide  mixed  with  air  from  the  copper 
sulphide  furnace,  is  passed.  The  hot  air  in  conjunction  with  iron  salts  in 
the  solution  oxidize  a  large  part  of  the  sulphurous  acid  to  sulphuric 

Method  of  Extracting  Copper  at  Rio  Tinto,  Spain.' — The  ore  treated  at 
Rio  Tinto  is  massive  iron  pyrite  containing  up  to  3  per  cent,  of  copper, 
which  has  been  disseminated  through  the  mass  by  a  secondary  en- 

The  well-known  method  adopted  for  the  extraction  of  the  copper,  and 
in  use  at  the  present  time,  consists  in  allowing  huge  heaps  of  the  mineral 
to  oxidize  under  the  influence  of  moisture  and  air,  and  subsequently 
washing  out  the  copper  sulphate  as  soon  as  it  is  formed,  by  running  water 
through  the  heap. 

The  application  of  this  system  depends  largely  on  the  state  in  which 
the  copper  occurs  in  the  mineral.  If  it  exists  as  chalcopyrite — CiiFeS2 — 
the  copper  will  not  oxidize  by  simple  exposure  to  the  air,  in  one  case  it 
having  taken  many  years  to  oxidize  10  per  cent,  of  the  copper  originally 
present  in  the  ore.  If  the  copper  is  in  the  form  of  CuS,  the  oxidation 
proceeds  very  slowly.  The  best  form  for  solution  is  CujS,  or  copper 
glance,  which  constitutes  the  bulk  of  the  copper  in  Rio  Tinto  pyrite. 
These  statements  may  be  made  clearer  from  the  following  account  of 
the  reactions  that  take  place  during  the  oxidation. 

When  the  mineral  is  exposed  to  free  access  of  air  and  moisture,  some 
ferrous  sulphate  is  formed  in  accordance  with  the  following  reaction: 

(1)  FeS2  +  70  +  H2O=FeSO,-hH2SO4 

This  ferrous  sulphate  becomes  readily  oxidized  by  the  air  to  ferric 

(2)  2FeSO,  +  H2S04  +  0=re2(SOj3  +  H,0, 

and  it  is  due  to  the  reaction  of  this  ferric  sulphate  on  the  copper  sulphides 
that  the  copper  is  rendered  soluble,  as  is  shown  by  the  following  chemical 

(3)  Fe,(SO,)3  +  Cu2S  =  CuS04-F2FeSO^  +  CuS. 

(4)  Fe,(SOj3  +  CuS  +  03  +  H,0  =  CuSO,  +  2FeSO,  +  H2SO, 

The  reaction  No.' 3  takes  place  fairly  rapidly  and  causes  half  the  copper 
to  go  into  solution  within  a  few  months,  while  reaction  No.  4  proceeds 

'C.  H.  Jones,  Trans.  Am.  Inst.  Min.  Engs.,  Vol.  XXXV,  1905. 


much  more  slowly,  and  requires,  under  the  most  favorable  conditions, 
about  two  years  to  extract  80  per  cent,  of  the  remaining  half  of  the 

In  the  laboratory  at  Rio  Tinto  a  method  has  been  worked  out  to 
determine  the  state  of  combination  of  the  copper  in  any  particular  min- 
eral. This  method  depends  on  the  action  of  the  mineral  on  various 
solutions  under  constant  conditions  of  dilution  and  temperature,  and, 
though  necessarily  somewhat  arbitrary,  it  shows  with  considerable 
accuracy  the  form  of  the  sulphide  in  which  the  copper  exists,  and  conse- 
quently whether  the  copper  can  be  readily  extracted  by  washing  in 

In  practice  the  method  adopted  to  bring  about  the  desired  oxidation 
is  as  follows:  A  site  is  chosen  for  the  formation  of  the  heap  where  the 
ground  is  sufficiently  concave  and  sloping  to  enable  the  copper  liquor 
that  is  formed  to  collect  and  run  out  at  the  base  of  the  heap.  On  the 
ground  is  first  arranged  a  net  work  of  air  flues,  made  of  rough  stones  and 
having  an  internal  diameter  of  12  in.  Vertical  chimneys,  50  ft.  distant 
from  one  another,  are  built  in  the  same  manner  and  connect  with  the 
ground  flues.  Care  is  taken  that  the  mouths  of  the  ground  flues  are  kept 
open  and  not  covered  by  ore.  The  mineral,  the  lump  portion  of  which 
has  been  passed  through  jaw  breakers  to  be  reduced  to  pieces  not  larger 
than  from  2  to  3  in.  across,  is  now  tipped  from  side-tip  wagons  at  the 
highest  part  of  the  selected  side  over  and  around  the  stone  flues.  Lump 
and  fines  are  alternately  dumped  until  the  height  of  the  mass  at  the  edge 
is  about  30  ft.,  the  upper  surface  of  the  mineral  being  kept  level.  A  heap 
of  this  form  approximately  contains  100,000  tons  of  ore. 

As  the  mineral  is  added,  the  building  of  the  stone  chimneys  keeps  pace, 
in  order  to  have  a  clear  opening  to  the  top  of  the  heap.  The  surface  of 
the  heap  is  formed  into  squares  by  means  of  ridges  of  the  mineral,  the 
size  of  these  squares  depending  on  the  porosity  of  the  heap.  The  func- 
tion of  these  ridges  is  to  enable  the  water  to  be  run  on  locally  over  the 
surface  of  the  heap  in  order  to  insure  that  all  parts  are  equally  washed 
and  that  the  water  does  not  run  through  the  heaps  in  channels.  A 
system  of  gutters  is  also  arranged  so  that  the  water  can  be  run  on  to  all 
parts  of  the  mass.  As  the  heap  is  being  formed,  water  is  run  on  and  the 
copper  sulphate  existing  in  the  mineral  is  extracted;  also,  the  water 
provides  the  moisture  for  the  oxidation  to  take  place  in  accordance  with 
the  equations  given  above.  The  mineral  in  the  heap  is  then  allowed  to 
oxidize,  which  it  does  pretty  rapidly,  as  evidenced  by  the  heat  produced, 
the  temperature  in  the  chimneys  rising  to  from  170  to  180°  F.  As  the 
temperature  increases,  the  surface  openings  of  the  chimney  may  be 
closed  in  order  to  allow  the  oxidation  to  spread  through  the  heap.  The 
surface  gradually  shows  a  brownish  coloration,  due  to  the  dehydration  of 
the  buff-colored  basic  ferric  salt  that  forms  on  top  of  the  mass,  and  its 


gradual  heating  up  may  be  noted  by  this  drying  action.  The  greatest 
care  must  be  taken  not  to  allow  the  heaps  to  fire,  for  if  once  started  it  is 
very  difficult  to  extinguish.  When  the  oxidation  has  proceeded  as  far  as 
it  is  safe  to  allow  it,  water  is  run  on  at  the  rate  of  about  50  cubic  meters 
per  hour,  until  the  soluble  copper  is  leached  out.  The  heap  is  then  allowed 
to  reoxidize  and  the  washing  is  repeated.  After  about  a  year  has  elapsed 
the  surface  requires  "refilling,"  and  the  squares  are  rearranged  so  that 
the  places  where  the  ridges  were  before  are  now  the  middle  of  the  squares. 
The  gutters  also  are  shifted.  At  the  edge  of  the  heap  for  a  distance  of 
some  yards  the  mineral,  which  has  become  cemented,  holds  a  consider- 
able quantity  of  copper  salts  and  is  dug  down  into  terraces  in  order  that 
this  copper  may  be  extracted  by  washing. 

When  the  copper  is  reduced  to  0.3  per  cent,  the  heap  is  considered 
washed  and  the  mineral,  containing  49.5  per  cent,  of  sulphur,  is  removed 
and  exported  as  "washed  sulphur  ore"  and  used  for  the  manufacture  of 
sulphuric  acid. 

Successful  heap  washing  depends  on  the  efficient  ventilation  of  the 
mass,  the  trouble  usually  being  a  too  great  excess  of  fines  produced  in 
mining  the  ore,  which  cement  hard  and  clog  up  the  air  passages. 

The  copper  liquor  as  it  runs  from  the  heap  contains  some  ferric  iron 
in  solution,which  as  will  be  shown  later,  is  very  objectionable.  In  order 
to  remove  the  ferric  iron  the  liquor  is  run  over  a  smaller  heap  of  fresh 
mineral  known  as  a  "filter  bed,"  which  reduces  the  ferric  iron.  This 
"bed"  is  laid  inside  a  reservoir  formed  by  a  masonry  dam  across  a  small 
ravine,  and  the  liquor  after  percolating  through  the  mineral  remains  in 
contact  with  it  until  it  is  required  to  be  drawn  off  to  the  precipitating 
tanks.  When  the  mineral  is  fresh  the  reduction  of  the  ferric  iron  takes 
place  rapidly,  due  to  the  CujS,  as  shown  by  equation  3,  but  the  iron 
pyrite  itself  has  an  effective  reducing  action  on  ferric  iron  in  solution 
according  to  the  following  equation: 

(5)  7Fe2(SOJ3+FeS2  +  8H20  =  15FeSO,+8H2SO,. 

The  principal  constituents  of  the  liquor  as  it  enters  the  cementation 
tanks  are  as  follows,  the  figures  given  representing  the  grams  per  cubic 
meter  or  units  per  million  parts:  Copper  4000,  ferric  iron  1000,  ferrous 
iron  20,000,  free  sulphuric  acid  10,000,  and  arsenic  300.  The  large 
quantities  of  ferrous  iron  and  free  sulphuric  acid  present  are  due  to  the 
fact  that  the  waste  liquor  from  the  cementation  tanks  after  the  copper 
has  been  precipitated,  is  pumped  back  and  used  for  washing  the  heaps  in 
addition  to  fresh  water,  and  consequently  these  solutions  tend  to  become 
concentrated.  The  liquor  is  then  run  from  the  reservoirs  at  about  300 
cubic  meters  per  hour  through  the  precipitation  tanks  over  pig  iron  in 
order  to  precipitate  the  copper  in  the  form  of  "cement  copper."  These 
cementation  tanks  are  arranged  in  series  on  the  slope  of  a  hill,  the  liquor 





passing  backward  and  forward  until  it  is  discharged  from  the  lowest  tank 
qf  the  series  free  from  copper. 

Each  series  consists  of  three  tanks  in  parallel  arranged  so  that  the 
liquor  can  be  divided  and  passed  along  as  many  tanks  as  necessary, 
depending  on  the  quantity  of  liquor  that  is  being  run  through  and  on  the 
varying  temperature  of  the  liquor  with  different  seasons,  the  hotter  the 
solution,  which  in  summer  reaches  100°  F.,  the  faster  the  rate  of  precipita- 
tion. Each  tank  is  about  320  ft.  long,  5.5  ft.  wide  by  2.2.5  ft.  deep,  and 
has  a  slope  varying  from  2  per  1000  in  the  first  series  to  11  per  1000  in 
the  last,  the  reason  for  the  increase  in  slope  being,  that  as  the  liquor 
becomes  impoverished  in  copper  the  free  acid  present  is  more  active  in 
wastefully   dissolving   the   pig   iron — an   action   which   is   considerably 

Fig.  47. — Rio  Tinto  leaching  plant,  Spain.     General  \'iew  of  mineral  licaps,  copper 
liquor  dam,  and  precipitating  tanks. 

diminished  by  increasing  the  velocity  of  the  liquor  by  means  of  the  in- 
creased slope  of  the  tanks.  The  tanks  themselves  are  made  of  9X3  in. 
boards  attached  to  wooden  frames  set  in  cement,  the  space  between 
parallel  tanks  being  filled  with  stone  and  cement,  constituting  a  wall 
.supporting  the  sides  of  the  tanks.  Fig.  46  shows  the  method  of  removing 
the  cement  copper  and  Fig.  47  a  general  view  of  the  mineral  heaps, 
copper  liquor  dam  and  precipitating  tanks. 

No  metal  is  used  in  the  tank  construction,  hard  wood  pegs  being  em- 
ployed to  attach  the  boards  to  the  frames.     The  spaces  between  the  boards 



are  carefully  caulked  with  oakum  and  pitch  in  order  to  render  the  tank 
water-tight.  At  each  end  of  the  tank  is  an  arrangement  by  which  a  door 
can  be  dropped  in  and  luted  so  as  to  cut  out  that  particular  tank,  and  there 
are  also  wooden  plugs  that  can  be  removed  so  that  the  liquor  from  that 
tank  can  be  run  off,  thus  allowing  for  the  removal  of  the  precipitated 
copper.  A  few  old  boards  are  placed  on  the  bottom  of  the  tanks  for 
their  protection  and  on  these  are  piled  up  the  pigs  of  iron  which  are  laid 
across  the  tank  at  the  bottom,  the  next  layer  being  at  right  angles  to  the 
first,  and  so  on  until  the  tank  is  filled;  each  foot-length  of  the  tank  con- 
tains about  one  ton  of  iron.  The  liquor  is  allowed  to  run  through  the 
system  of  tanks  and  needs  no  attention  except  to  remove  the  precipitated 
copper  and  to  add  fresh  iron.  The  "sahda"  liquor  containing  from  15 
to  20  grm.  of  copper  per  cubic  meter  is  allowed  to  run  to  waste,  for  the 
reason  that  about  this  copper  content  the  amount  of  iron  required  to 
precipitate  the  copper  equals  the  value  of  the  copper  recovered.  Daily 
some  of  the  tanks  are  cleaned  out  by  being  closed  as  described  above,  the 
liquor  meanwhile  passing  down  the  other  tanks  of  the  series;  the  liquor 
is  run  off  into  settling  tanks,  any  copper  in  suspension  being  recovered; 
all  of  the  iron  is  removed  from  the  tank  and  piled  on  to  the  dividing  wall, 
at  the  same  time  the  copper  adhering  to  the  iron  is  knocked  off  and  thrown 
back  into  the  tank.  The  dirty-looking  precipitate  is  then  transferred  to 
the  cleaning  and  concentrating  plant,  the  iron  is  replaced  in  the  tank 
and  the  liquor  again  allowed  to  run  through  it.  This  crude  precipitate, 
containing  about  70  per  cent,  copper,  is  thrown,  a  little  at  a  time,  on  a 
perforated  copper  plate  at  the  head  of  a  long  launder  or  tank  and  is 
washed  through  the  plate  by  a  strong  stream  of  water  from  a  small 
nozzle.  The  material  that  does  not  pass  through  the  screen  consists  of 
leaf-copper  and  small  pieces  of  iron;  this  material  is  thrown  into  a  heap 
and  afterward  sorted  over  by  girls  who  remove  the  pieces  of  iron.  The 
precipitate  that  passes  into  the  launder  is  repeatedly  turned  over  against 
the  stream  of  water  and  by  this  simple  means  a  concentration  is  effected. 
The  first  few  yards  of  the  launder  contain  a  red  precipitate  known  as 
"No.  1  precipitate",  containing  94  per  cent,  copper  and  less  than  0.3 
per  cent,  arsenic;  following  this  is  "No.  2  precipitate"  containing  92 
per  cent,  copper  and  between  0.3  per  cent,  and  0.75  per  cent,  arsenic, 
while  below  is  the  "No.  3  precipitate";  this  is  in  a  state  of  very  fine 
division  and  contains  on  an  average  50  per  cent,  copper  and  5  per  cent, 
arsenic.  This  last  named  portion,  which  carries  all  the  graphite  from 
the  pig  iron,  contains  the  bulk  of  the  antimony  and  bismuth  that  is  also 
precipitated  from  the  liquors.  Classes  No.  1  and  No.  2  are  removed  to  the 
drying  sheds  and  bagged  for  shipment  to  the  refinery;  the  No.  3  precipi- 
tate is  removed,  moistened  with  acid  liquors,  made  into  balls  by  hand  and 
dried  in  the  sun.  These  balls  become  cemented  hard  and  can  be  readily 
transferred  to  the  smelter,  where  they  form  part  of  the  charge  for  the 


blast  furnaces  and  are  run  down  to  matte  to  be  subsequently  bessermerized, 
thus  effectively  removing  the  arsenic,  antimony,  and  bismuth  they  con- 
tain. The  reactions  that  take  place  in  the  cementation  tanks  are  given  in 
equations  5,  6,  and  7. 

The  first  reaction  that  occurs  in  the  liquor  running  over  metallic  iron 
is  the  reduction  of  the  ferric  sulphate  to  ferrous  sulphate,  the  final  result 
being  in  accordance  with  the  equation: 

(5)  Fe2(SO,)3  +  Fe  =  3FeSO,. 

This  action  causes  the  consumption  of  the  pig  iron  without  any  cor- 
responding yield  in  copper,  and  consequently  should  be  avoided  as  far  as 
possible  by  having  all  the  iron  in  the  ferrous  condition.  The  second 
reaction  is  the  precipitation  of  the  metallic  copper,  brought  about  by 
galvanic  action.  The  iron  becomes  coated  with  copper,  and  thus  the  iron 
and  copper  in  the  acid  liquor  constitute  a  galvanic  couple  with  a  consider- 
able difference  of  potential.  It  is  due  to  the  electrolytic  action  that  the 
copper  and  all  other  metals  present  that  are  electro-negative  to  iron  will 
be  precipitated.  The  ultimate  action  of  the  precipitation  may  be  chemi- 
cally expressed  by  the  following  equation: 

(6)  CuS0,-|-Fe=FeS04-hCu. 

Besides  the  reactions  above  mentioned  there  is  one  which  causes  the 
liberation  of  hydrogen,  as  evidenced  by  the  bubbles  of  gas  that  may  be 
observed  to  arise  in  the  tank  liquor.  This  action,  Avhich  causes  a  waste- 
ful consumption  of  iron,  may  be  expressed  as  a  final  result  by  the  follow- 
ing equation: 

(7)  Fe  +  H,S0,=FeS0,  +  2H. 

These  three  equations  constitute  the  main  reactions  that  take  place  in  the 
precipitating  tanks. 

While  the  liquor  is  fairly  strong  in  copper,  the  copper  is  mostly 
precipitated  in  a  coherent  form,  but  in  the  later  stages,  as  the  liquor 
becomes  impoverished  it  is  precipitated  in  a  powdery  state — a  condition 
which  is  more  effective  in  its  galvanic  action  with  the  iron,  and  thus 
unfortunately  causes  a  larger  precipitation  of  arsenic  and  other  impuri- 
ties than  in  the  earlier  stages.  In  the  later  stages  also  the  "solution 
reaction, "  of  iron  and  sulphuric  acid  as  given  in  equation  No.  7  goes  on  to 
11  proportionately  greater  extent  than  does  the  precipitation  of  the  copper, 
and  consequently  the  cost  of  pig  iron  in  precipitating  the  copper  varies 
inversely  as  the  quantity  of  copper  in  the  liquor. 

By  keeping  careful  watch  on  the  reduction,  as  far  as  possible,  of  the 
ferric  iron  before  the  liquor  enters  the  tanks,  and  by  giving  it  sufficient 
velocity  through  the  tanks,  a  strongly  acid  liquor  such  as  given  above 
during  a  years'  working  will  not  consume  more  than  1.4  units  of  pig  iron 


(containing  92  per  cent,  iron)  to  1  unit  of  copper  precipitated.  A  valu- 
able check  on  the  iron  being  consumed  can  easily  be  kept  by  the  labor- 
atory, by  analyzing  the  liquors  before  entering  and  after  leaving  the 
tanks,  and  from  these  analyses  the  quantity  of  iron  that  is  being  con- 
sumed can  be  calculated. 

The  following  table^  gives  the  analysis  (in  grams  per  cubic  meter)  of 
the  entering  and  outgoing  liquors  of  the  precipitation  plant  at  Rio 




i    Leaving 



Ferrous  iron 

,      ...'            2,710 

.  .      ..i          13,908 





'■  1.26  to  1  Cu 



4,129  ' 







1.26  to  1  Cu 


Free  acid 

Total  solids 

Specific  gravity 




The  best  results  are  obtained  when  the  solution  is  slightly  acid,  as  it 
tends  to  accelerate  precipitation  and  prevents  a  falling  out  of  basic  salts 
of  iron  while  the  precipitation  is  going  on. 

At  Tharsis  the  inclination  of  the  precipitating  tanks  is  as  follows: 

For  the  first  40  per  cent,  of  the  copper,  1  in  200 

For  the  next  30  per  cent,  of  the  copper,  1  in  150 

For  the  next  20  per  cent,  of  the  copper,  1  in  100 

For  the  remainder,  1  in     50 

At  Rio  Tinto  and  Tharsis,  1.4  units  of  pig  iron  containing  92  per  cent, 
iron  are  consumed  on  an  average,  per  unit  of  copper  obtained.  Both*  at 
Rio  Tinto  and  Tharsis  the  liquor  traverses  about  3  kilometers  before  all 
the  copper  is  precipitated.  Sixty  per  cent,  of  the  copper  in  the  liquors 
is  precipitated  within  the  first  700  meters  of  the  tanks.  Over  70  per  cent, 
of  the  precipitate  contains  more  than  94  per  cent,  copper. 

The  following  points  have  been  established  by  the  practice  at  Rio 

1.  The  complete  analysis  of  the  solutions  both  before  and  after 
treatment  is  essential  to  prevent  undue  consumption  of  iron. 

2.  Free  acid  to  be  eliminated  as  much  as  possible. 

3.  Mechanical  contrivances  will,  in  a  measure,  overcome  these  diffi- 
culties. The  inclination  of  precipitating  tanks  and  consequent  velocity 
of  current  should  be  in  inverse  ratio  to  the  amount  of  copper  present; 
the  less  the  copper  and  the  greater  the  free  acid  and  ferric  iron,  the 
greater  inclination  necessary. 

'F.  H.  Probert,  Mining  and  Scientific  Press,  January  4,  1908. 


4.  As  large  a  surface  of  iron  should  be  exposed  to  the  liquors  as 

5.  Aeration  of  the  liquor  by  tumbling  through  the  air  is  objectionable, 
since  it  has  an  oxidizing  effect  and  so  increases  the  consumption  of  iron. 

6.  In  the  course  of  the  flow  of  the  liquor  through  the  precipitating 
tanks  there  is  a  place  where  the  cost  of  iron  exceeds  that  of  the  value  of 
the  copper  obtained. 

7.  The  warmer  the  liquors,  within  certain  limits,  the  faster  is  the 
precipitation  of  the  copper. 

M.  P.  Truchot,  chief  chemist,  Huelva,^  estimates  that  to  extract  cop- 
per by  natural  weathering  alone,  20  years  would  be  necessary;  roasting 
destroys  the  vegetation  of  the  country;  there  therefore  remains  the  wet 
methods  where  the  reactions  may  be  hastened  by  the  use  of  a  regulated 
supply  of  air  and  water.  In  preparing  the  heaps  for  oxidizing  leaching 
of  the  copper,  the  fines  amount  to  80  per  cent. ;  the  lumps,  20  per  cent. 
The  two  grades  are  placed  in  alternate  layers  and  the  top  of  the  heap 
finished  with  fines,  to  prevent  too  rapid  filtration.  Practice  varies  as  to 
the  temperature  of  the  heap,  which  may  be  as  high  as  82  to  90°  C.  where 
it  is  sought  to  promote  oxidation  and  to  increase  the  rate  of  leaching, 
while  in  other  cases,  the  temperature  is  not  allowed  to  exceed  30  to 
32°  C.  Since  the  higher  temperature  is  dangerous,  one  of  no  more  than 
45  to  60°  C.  is  recommended. 

It  takes  6  to  7  years  to  exhaust  one  of  these  heaps  of  100,000  tons;  the 
exhausted  ore  then  contains  0.25  to  0.30  per  cent,  of  copper.  The  more 
permeable  copper  schists  leach  more  rapidly,  however,  taking  but  three 
or  four  years.  While  the  oxidation  of  chalcopyrite  proceeds  but  slowly, 
the  rate  can  be  increased  by  finely  pulverizing  it  and  distributing  it 
throughout  the  pile. 

The  copper  solution  from  the  heaps  varies  from  0.015  to  0.5  or  even 
0.6  per  cent.  Cu.  It  is  of  a  reddish-green  color  containing  ferric  and 
ferrous  sulphates,  free  sulphuric  acid,  copper  sulphate,  and  other  salts. 

AVhen  the  water  supply  is  abundant,  and  the  free  acid  is  conse- 
quently low,  the  consumption  of  pig  iron  varies  from  1.3  to  1.5  tons  per 
ton  of  copper  precipitated.  This  seldom  occurs,  however,  and  generally 
there  is  needed  1.75  to  2  tons  per  ton  of  copper  extracted;  that  is  more 
than  double  the  theoretical  quantity. 

Treatment  by  Heap  Roasting  and  Leaching. — Instead  of  extracting 
or  treating  all  the  material  by  weathering,  the  extraction  of  the  copper 
may  be  expedited  by  first  giving  the  heaps  a  slow  roast,  by  which  process 
much  of  the  copper  in  the  cupriferous  pyrites  is  converted  into  sulphate 
and  can  readily  be  leached.  The  production  of  sulphate  of  copper  l3y 
slow  heap  roasting  can  only  be  used  on  ores  that  contain  proportionately 
large  amounts  of  iron  pyrites  and  small  amounts  of  copper  p3rrites  or 
'  L'  Echo  des  Mines  et  de  la  Metallurgie,  1906,  p.  482. 


other  copper  sulphides.  In  the  slow  and  imperfect  roasting  in  heaps  or 
stalls  not  all  of  the  copper  sulphide  will  be  converted  into  sulphate; 
a  part  will  always  remain  unaffected,  and  another  part  will  be  trans- 
formed into  oxides.  In  order  to  extract  as  much  copper  as  possible  from 
the  ores,  after  the  roasting  has  been  finished  and  the  sulphate  of  copper 
leached  out,  the  ores  are  allowed  to  weather,  as  for  the  fresh  or  unroasted 
ores,  already  described,  whereby  in  the  course  of  time  the  copper  is  con- 
verted into  sulphate,  by  natural  weathering. 

The  production  of  sulphate  of  copper  by  roasting,  followed  by  weath- 
ering of  the  leached  roasted  ore,  was  used  for  a  long  time  at  Rio  Tinto, 
in  Spain.  According  to  Schnabel,'  the  ores  thus  treated  were  cuprifer- 
ous pyrites  with  1 1/2  to  2  percent,  copper;  these  were  slowly  roasted  in 
heaps  of  200  to  1500  tons  upon  a  bed  of  brushwood,  firewood  or  coals. 
The  200  ton  heaps  were  hemispherical,  26  ft.  in  diameter  at  the  base  and 
11  ft.  6  in.  high.  The  larger,  1500  ton  heaps  were  elliptical  in  plan,  the 
longer  axis  of  the  ellipse  being  55  ft.  9  in.,  and  the  shorter  axis  32  ft. 
10 'in.;  their  height  was  also  11  ft.  6  in.  Air  was  admitted  by  means  of  a 
system  of  channels  traversing  the  heaps.  The  smaller  heaps  burned  for 
two  months,  the  larger  ones  for  six.  To  roast  100  tons  of  ore  required  in 
the  small  heaps  27  cu.  ft.  of  wood  and  in  the  larger  ones  9  cu.  ft.  The 
yield  of  copper  was  greater  in  the  small  heaps  than  in  the  larger  ones. 
The  roasted  ores  were  leached  for  50  hours  by  -which  means  the  copper 
present  as  sulphate  was  washed  out  of  them.  The  exhausted  residues 
still  contained  0.4  to  0.5  per  cent,  of  copper,  chiefly  as  sulphide,  to  extract 
which  the  ores  were  allowed  to  weather.  With  this  object  they  were  piled 
on  a  system  of  horizontal  flues  built  of  dry  stone,  that  air  could  circulate 
through  the  pile  of  ore.  The  vertical  flues  were  continued  in  proportion  as 
the  heaps  got  higher  by  the  piling  on  of  additional  ore.  As  soon  as  the 
damp  heap  had  reached  a  certain  height,  the  sulphides  began  to  de- 
compose, as  was  shown  by  the  rising  temperature.  By  checking  the  air 
supply  it  was  kept,  if  possible,  from  rising  so  high  that  the  heap  took  fire. 
From  time  to  time,  the  heap,  or  a  portion  of  it,  if  it  was  a  large  one,  was 
leached  out,  and  the  liquor  conducted  to  the  precipitating  tanks.  The  ex- 
haustion of  the  heaps,  which  were  continually  being  increased,  and  which 
may  reach  500,000  tons,  will  not  be  completed  in  measurable  time,  as,  in 
spite  of  frequent  leaching,  the  weathering  proceeds  very  slowly.  It  is 
even  held  that  these  huge  heaps  will  still  be  producing  sulphate  of 
copper  long  after  the  mines  shall  have  been  worked  out. 

The  cost  of  producing  copper  at  Rio  Tinto,  Spain,  by  heap  roasting 
and  leaching,  is  stated  to  be  about  11.55  per  ton,  divided  as  follows: 
Mining  89  cents,  roasting  about  18  cents,  including  labor  in  building 
heaps,  etc.,  precipitating  and  collecting  about  56.6  cents.  Sixty-six  tons 
of  ore  were  required  to  produce  one  ton  of  metallic  copper. 

'  Handbook  of  Metallurgy,  Vol.  I,  p.  212. 



The  cost  of  producing  one  ton  of  metallic  copper  by  the  cementation 
process  as  carried  out  there  is  about  $144.00. 

The  first  cement  copper  was  produced  in  Spain  at  Rio  Tinto,  in  1752 
from  heaps  of  low-grade  sulphide  ore  that  had  undergone  decomposition 
through  natural  processes.  The  copper  was  leached  out  by  water,  the 
metal  being  precipitated  on  iron.  It  was  first  thought  to  be  merely  a 
coating  of  copper  on  the  iron,  but  it  was  found  if  left  long  enough,  the 
replacement  became  complete. 


Atmospheric  Oxidation,  Without  Burning.^ — In  atmospheric  oxidation 
of  cupriferous  pyrites,  and  subsequent  extraction  of  copper  by  leaching, 
as  carried  on  in  Portugal  and  Spain,  arsenic  and  antimony  are  to  some 
extent  dissolved  and  precipitated  with  the  copper.  The  proportion  in 
which  they  are  separated  and  precipitated  is  only  a  fraction  of  the 
amount  contained  in  the  ores,  as  shown  by  the  following  tables  by 
Allen  Gibb : 



Per  cent, 

Per  cent,  relative 

Cu  =  100per 


Per  cent, 

Per  cent,  relative 

Cu-100  per 


Total  per  cent, 
of  elimination 



20  0 





Arsenic ... 




Burning  and  Subsequent  WashiTig. — In  this  process  largely  used  in 
Spain  and  to  some  extent  in  Portugal,  for  the  extraction  of  copper  from 
cupriferous  iron  pyrites,  there  is  a  considerable  elimination  of  arsenic, 
antimony,  and  bismuth  in  burning.  The  following  table,  by  Gibb,  may 
be  taken  as  fairly  typical  of  heap  roasting,  preparatory  to  leaching,  as 
conducted  in  Spain. 

Raw  Pyrites 

Burnt  Pyrites 

Per  cent, 

I  Per  cent,  relative  ' 
j  Cu  =  100per  : 
I  cent. 

Per  cent, 

Per  cent,  relative 

Cu  =  100  per 


Total  per  cent, 
of  elimination 

'  Allen  Gibb,  Trans.  A.  I.  M.  E.,  Vol.  XXXIII,  p.  667. 



o  45 

100  0 

3.32          ^ 





















The  chloride  processes  have  been  widely  applied  to  the  extraction  of 
copper  from  its  various  ores.  Hydrochloric  acid,  ferric  and  ferrous 
chlorides,  are  the  solvents  usually  employed. 

Hydrochloric  acid  presents  certain  advantages  over  sulphuric  acid  in 
the  technical  operation,  but  is  usually  more  expensive,  especially  in 
copper  mining  districts.  Sulphuric  acid,  in  the  process  of  operation, 
forms  ferric  sulphate,  which  when  exposed  to  the  air  is  decomposed  into 
basic  siilphate  of  iron  and  free  sulphuric  acid,  consequently  consum- 
ing more  iron  in  the  precipitation  of  the  copper.  Hydrochloric  acid  is 
less  apt  to  form  basic  salts  and  therefore  yield  solutions  that  contain  but 
little  free  acid,  and  which,  accordingly,  require  less  iron  for  the  precipita- 
tion of  the  copper,  than  do  solutions  containing  ferric  sulphate.  On 
the  other  hand  it  attacks  oxide  of  iron  more  energetically  than  does 
sulphuric  acid,  but  the  ferric  or  ferrous  chloride  so  formed  is  more  likely 
to  be  precipitated  out  of  the  solution  as  the  insoluble  ferric  oxide,  than 
from  a  sulphate  solution. 

Ordinarily  only  oxidized  ores  are  applicable  to  treatment  by  a  chloride 
process.     The  copper  may  be  dissolved  either  by 

Hydrochloric  acid,  or 
Metal  chloride. 

The  acid,  however,  is  the  solvent  usually  employed.  The  chlorides 
have  been  used  as  solvents  on  both  oxide  and  sulphide  ores  but  on 
sulphide  ores  the  action  has  been  too  slow  for  profitable  application, 
although  both  ferric  and  ferrous  chlorides  were  quite  extensively  used  at 
one  time. 

Hydrochloric  Acid. — If  iron  is  used  as  the  precipitant,  as  it  usually 
is,  the  hydrochloric  acid  process  consists  essentially  of  applying  dilute 
hydrochloric  acid  to  the  oxidized  ores  of  copper,  which  reacts  with  the 
copper  oxide  thus: 

CuO+2HCl  =  CuCl2  +  H20 

to  form  cupric  chloride.     Some  cuprous  chloride  may  also  be  formed: 
The  cupric  chloride  thus  formed  is  filtered  from  the  ore  and  pre- 
cipitated with  iron,  thus: 

CuCl2+Fe=FeCl2  +  Cu 

the  iron  and  copper  changing  places.  The  copper  is  precipitated  while 
the  iron  goes  into  solution  as  ferrous  chloride.  Theoretically  the  same 
amount  of  iron  is  required  to  precipitate  a  pound  of  copper  from  cupric 
chloride  solutions  as  from  cupric  sulphate  solutions  i.e.,  56  lb.  of  iron  to 
precipitate  63.6  lb.  of  copper.  In  practice,  however,  it  will  usually  take 
less  for  chloride  than  for  sulphate  solutions.  If  the  copper  in  solution 
is  in  the  form  of  cuprous  chloride,  only  half  the  amount  of  iron  is  required 


as  that  used  when  the  copper  is  in  the  form  of  either  cupric  chloride  or 
cupric  sulphate. 

If  hydrochloric  acid  is  used  as  the  solvent  is  takes  theoretically 
approximately  0.6  lb.  of  acid  to  extract  1  lb.  of  copper  as  cuprous  chloride, 
and  1.1  lb.  as  cupric  chloride. 

In  practical  operation,  much  more  acid  is  required,  the  amount 
depending  largely  on  the  nature  of  the  ore.  The  theoretical  amount  of 
acid  given  to  react  with  the  copper,  is  supposed  to  be  the  pure  acid,  free 
from  water.  Commercial  hydrochloric  acid,  is  the  HCl  dissolved  in 
water,  so  that  in  estimating  the  commercial  acid  required,  it  will  be 
necessary  to  first  know  the  acid  content,  which  can  be  determined  from 
its  specific  gravity. 

Concentrated  HCl  +  Aq  loses  HCl,  and  dilute  HCl  +  Aq  loses  water 
on  warming,  until  an  acid  of  constant  composition  is  formed,  containing 
20.18  per  cent.  HCl  with  a  specific  gravity  of  1.101  at  15°  C,  which  can 
be  distilled  unchanged  at  110°  C.  Concentrated  HCl+Aq  gradually 
gives  off  HCl  on  the  air  until  it  has  a  specific  gravity  of  1.128  at  15°  C, 
and  contains  15.2  per  cent.  HCl.  In  leaching  copper  ores,  only  very 
dilute  solutions  of  hydrochloric  acid  are  usually  employed,  rarely  exceed- 
ing 5  per  cent.  HCl. 

In  Stadtberg,  Westphalia,  hydrochloric  acid  was  formerly  used  to 
extract  copper  from  ores  containing  from  1  to  2  per  cent,  copper.'  The 
leaching  vessels  were  rectangular  tanks  of  wood  4  ft.  11/2  in.  high, 
packed  in  a  lied  of  clay  1  ft.  3  in.  thick,  then  fitted  with  a  grating  of 
wooden  bars,  upon  which  the  ore  was  piled.  The  first  tanks  held  29  tons 
of  ore,  but  those  erected  afterward  had  a  capacity  of  90  tons.  All  the 
tanks  were  situated  on  a  level.  The  leaching  was  methodical;  ores 
nearly  free  from  copper  being  treated  with  fresh  hydrochloric  acid  of  12 
to  13°  B.,  as  obtained  from  soda  works,  while  frejh  ore  was  treated 
with  partially  saturated  solution  until  the  latter  was  fully  saturated, 
which  took  place  at  19  to  20°  B.  The  various  solutions  were  allowed 
to  remain  12  hours  in  each  tank,  the  saturation  point  being  reached  in 
10  to  12  days.  The  solution  was  circulated  by  means  of  pumps  and 
bucket  wheels.  After  the  solution  had  percolated  through  the  ores,  it 
ran  out  thxough  a  plug  hole  to  which  the  bottom  of  the  tank  was  in- 
clined, into  receivers,  whence  it  was  again  lifted  to  its  proper  tank. 
The  exhausted  ore  was  allowed  to  lie  for  another  12  to  15  hours  in  water 
and  was  then  washed  for  12  hours  more.  Fresh  acid  was  diluted  with  a 
portion  of  the  mother-liquor.  For  each  100  parts  by  weight  of  copper, 
550  to  700  parts  of  hydrochloric  acid  12  to  15°  B.,  were  employed. 
This  process  replaced  the  sulphuric  acid  leaching,  previously  employed 
but  was  later  abandoned  when  the  carbonates  of  the  ore  were  replaced 
with  sulphides  in  depth. 

'Schnabel  Handbook  of  Metallurgy,  Vol.  I,  p.  200. 


At  Twiste,  in  Waldeck,  attempts  were  made  to  leach  malachite  and 
azurite  copper  ores  occurring  in  the  Bunter  Sandstein  (Lower  Trias)  and 
containing  1/2  to  1  per  cent,  of  copper,  by  means  of  hydrochloric  acid, 
but  had  to  be  abandoned  because  the  ores  contained  from  1/2  to  1  per 
cent,  of  lime  which  was  dissolved  by  the  acid  before  it  attacked  the  copper. 

Ferric  Chloride,  FeClj.— Ferric  chloride,  like  ferric  sulphate,  has  the 
property  of  dissolving  copper  from  its  oxide,  carbonate,  and  sulphide 
combinations.  If  the  sulphides  of  iron,  copper,  lead,  zinc,  arsenic  and 
antimony  are  treated  with  a  hot  solution  of  ferric  chloride,  the  minerals 
are  more  or  less  decomposed,  and  the  respective  metals  go  into  solution 
as  chlorides,  usually  with  the  liberation  of  sulphur.  Hydrochloric  or 
sulphuric  acid,  added  to  the  solution,  aids  the  reactions.  Silver  is  con- 
verted into  chloride  by  the  action  of  ferric  chloride.  Gold  remains 

If  ferric  chloride  is  used  as  a  solvent  for  ore  containing  the  copper  as 
cupric  sulphide  (CuS) ,  cupric  chloride  is  produced  and  the  ferric  chloride 
is  reduced  to  the  ferrous  chloride: 

CuS  +  2FeCl3  =  CuClj  +  2FeCl ,  +  S. 

If  the  copper  mineral  contains  the  copper  as  cuprous  sulphide  the 
reaction  will  result  in  the  formation  of  cupric  and  cuprous  chlorides: 

Cu,S  +  2FeCl3  =  2CuCl  +  2FeCl2  +  S 
Cu,S  +  4FeCl3  =  2CuCl2  +  4FeCl2  +  S. 

With  a  neutral  solution  of  ferric  chloride,  the  reaction  with  copper 
oxide  is: 

3CuO  +2FeCl3  =  SCuCl^  +Fe203 

although  some  cuprous  chloride  may  also  be  formed  by  the  reduction  of 
the  ferric  to  the  ferrous  chloride,  and  the  subsequent  reaction  of  the 
ferrous  chloride  on  the  copper  oxide.  If  the  solution  is  somewhat  acid, 
the  ferric  chloride  will  be  reduced,  but  the  ferrous  chloride  will  not  be 
precipitated  to  any  great  extent.  With  copper  carbonates  the  reaction 
is  much  the  same  as  for  oxides,  except  that  carbon  dioxide  is  liberated 
in  the  reaction. 

In  reacting  with  sulphide  ores,  the  ferric  chloride  is  reduced  to  the 
ferrous  chloride,  but  the  ferrous  chloride  does  not  further  react  with 
the  copper  sulphides.  In  any  event,  the  sum  total  of  the  reactions  may 
be  considered  to  be  the  reduction  of  the  ferric  to  the  ferrous  chloride, 
with  the  formation  of  cupric  and  cuprous  chlorides. 

The  copper  may  be  precipitated  with  iron: 

CuJlj  +Fe  =Cu  +FeCl2, 
2CuCl+Fe  =2Cu+FeCl2, 


ferrous  chloride  being  formed.  If  ferric  chloride  exists  in  the  solution 
during  precipitation  with  iron,  it  will  be  reduced  to  ferrous  chloride  at  the 
expense  of  the  iron. 

The  ferrous  chloride,  after  precipitation  of  the  copper,  may  be  re- 
generated back  to  ferric  chloride  by  the  action  of  air,  chlorine,  or  hydro- 
chloric acid.  If  the  neutral  ferrous  chloride  is  regenerated  by  agitation 
with  air,  some  of  the  iron  is  brought  into  the  condition  of  basic  salts,  and 
much  is  precipitated  as  the  ferric  oxide. 

6FeCl2  +  O3  =4FeCl3  +Fe203. 

From  this  equation  it  is  apparent  that  one-third  of  the  iron  is  pre- 
cipitated from  the  solution  as  the  insoluble  ferric  oxide  in  order  to  raise 
the  remaining  two-thirds  to  the  ferric  condition.  In  addition  to  the 
precipitated  feriic  oxide  there  may  be  formed,  in  netural  solutions,  in- 
soluble oxychlorides,  and  these  necessitate  a  further  loss,  both  of  iron  and 
chlorine.  If,  however,  iron  is  used  as  the  precipitant,  large  amounts  of 
ferrous  chloride  are  produced  which  may  then  be  brought  to  the  ferric 
condition  for  reuse  as  a  solvent,  and  the  loss  of  iron  as  ferric  oxide  or 
oxychloride,  is  hot  a  serious  matter;  in  fact  an  elimination  of  a  certain 
amount  of  the  iron  is  an  absolute  necessity. 

If  the  ferrous  chloride  solution  is  regenerated  with  chlorine,  oxy- 
chlorides are  not  formed,  although  some  iron  may  be  precipitated  as  fer- 
ric oxide.  The  chlorine  may  be  produced  either  chemically  or  elec- 
trolytically,  by  any  of  the  well-known  methods.  The  chlorine  combines 
directly  with  the  ferrous   chloride  to  produce  the  ferric   chloride; 

FeCl2  +  Cl=FeCl3. 

The  solubility  of  copper  from  sulphide  ores  with  ferric  chloride  solu- 
tions, depends  much  on  the  way  the  copper  is  mineralogically  combined. 
The  copper  in  the  form  of  chalcocite  is  much  more  readily  soluble  than 
in  the  form  of  chalcopyrite,  while  gray  copper  remains  quite  unaffected. 
Careful  roasting  at  a  low  temperature  makes  the  copper,  in  any  of  its 
sulphide  combinations,  readily  soluble,  but  when  roasted  at  a  high 
temperature,  the  copper  is  quite  insoluble  either  in  acids  or  solutions  of 
ferric  salts.  This  is  doubtless  due  to  the  formation  of  silicates  and  of 

Doetsch  Process. — Ferric  chloride  acts  on  cupric  and  cuprous  sul- 
phides to  form  cupric  and  cuprous  chlorides,  while  the  ferric  chloride  is 
reduced  to  the  ferrous  condition.     The  reactions  may  be  expressed  thus: 

CuS  +2FeCl3=  CuCl2  +  2FeCl2  +  S 
Cu2S+2FeCl3  =  2CuCl  +2FeCl2  +  S. 

On  these  reactions  are  based  the  Doetsch  Process,  formerly  used  ex- 
tensively at  Kio  Tinto  in  Spain.     The  ore  there  treated  contained  on  an 


average  of  2.7  per  cent,  copper.  The  copper  passed  into  solution  while 
the  pyrite  was  practically  unaffected  in  the  leaching  operation.  The 
process  has  the  advantage  of  reducing  the  waste  of  iron  in  the  precipitating 
tanks  by  avoiding  the  formation  of  ferric  sulphate.  In  this  way  one  ton 
of  copper  precipitate  was  obtained  with  an  expenditure  of  about  an 
equal  weight  of  iron. 

In  the  process  as  carried  out  at  Rio  Tinto  the  ore  was  crushed  to 
about  1/2  in.,  mixed  with  0.5  per  cent,  common  salt  and  a  like  amount  of 
ferrous  sulphate,  and  built  into  large  heaps.  These  heaps  were  from  12 
to  16  ft.  high  and  50  ft.  square  at  the  base.  A  solution  of  ferric  chloride 
was  run  in  a  continuous  steam  upon  these  heaps. 

It  took  about  four  months  to  extract  1.34  per  cent,  of  the  copper,  or 
50  per  cent,  of  the  total  copper  content.  After  two  years  2.20  per  cent, 
was  extracted  or  80  per  cent,  of  the  contained  copper.  The  final  loss 
was  about  0.48  per  cent. 

The  solutions  leached  through  the  heaps,  containing  cupric  and  cup- 
rous chlorides  as  also  ferric  and  ferrous  chlorides,  were  precipitated  with 

CuCl2+Fe  =  Cu   -hFeClj 
2CuCl+Fe  =  2Cu+FeCl2 

ferruos  chloride  being  formed.  The  ferric  chloride,  by  its  action  on  the 
ore,  is  changed  to  ferrous  chloride;  small  amounts  of  ferric  chloride 
may  remain  unchanged  in  the  solution,  but  this  reacts  with  the  iron  in  the 
precipitating  tanks  and  may  be  reduced  there  to  the  ferrous  condition. 

To  make  the  process  continuous,  the  ferrous  chloride  is  regenerated 
back  to  ferric  chloride,  which  is  accomplished  by  bringing  the  ferrous 
chloride  solution  in  contact  with  chlorine,  in  scrubbing  towers. 

The  solution  which  ran  from  the  ore  heaps  contained  5  to  7  kilogrm. 
of  copper  per  1000  kilogrm.,  or  1  cubic  meter. 

The  chlorine  gas,  for  the  regeneration  of  the  ferrous  chloride  to  ferric 
chloride,  was  produced  by  heating  salt  with  ferrous  sulphate  in  a  re- 
verberatory  furnace,  holding  500  lb.  An  abundance  of  air  was  admitted 
through  the  working  doors.  The  ferrous  sulphate  reacting  with  salt  in 
the  presence  of  air,  produces  chlorine,  sodium  sulphate,  and  ferric  oxide. 

2FeSO,  +  4NaCl  +  03=Fe203  +  2Na2SO,  +  4Cl. 

The  ferrous  sulphate  used  is  found  in  large  quantities  on  the  shores 
of  the  Rio  Tinto  river.  The  immense  heaps  of  low-grade  ore,  "Toreros" 
being  leached  by  natural  cementation,  also  furnish  salts  of  ferric  and 
ferrous  sulphate.  During  the  reaction  in  the  reverberatory  furnace, 
some  hydrochloric  acid  is  formed;  this  is  converted  into  chlorine  by  the 
action  of  manganese  dioxide  of  which  there  was  a  certain  amount  placed 
near  the  flues  of  the  furnace. 


The  gases  from  the  furnace,  consisting  largely  of  chlorine  with  pos- 
sibly some  hydrochloric  acid,  were  conducted  to  scrubbing  towers,  where 
they  came  in  contact  with  the  ferrous  chloride  solution  from  the  precipi- 
tating tanks,  and  the  ferrous  chloride  reconverted  to  the  ferric  chloride. 

The  Doetsch  process  was  used  both  for  raw  and  roasted  ore.  When 
roasted,  the  roasting  was  performed  in  heaps  of  truncated  pyramidal 
form,  10  ft.  high,  with  a  base  20x26.5  ft.  for  those  containing  800  tons, 
and  30X26.6  ft.  for  those  containing  1200  tons. 

At  the  bottom  of  the  heaps,  one  transverse  and  three  longitudinal 
flues,  about  20  in.  square,  were  formed  by  the  larger  blocks  of  mineral. 
Those  were  the  firing  passages,  and  communicated  with  vertical  chimneys, 
of  which  there  were  two  in  the  800-ton,  and  three  in  the  1200-ton  heaps. 
The  mass  of  the  heaps  was  made  of  lumps  a  little  above  nut  size.  Salt 
was  added  in  the  proporiton  of  14  tons  to  800  tons  of  pyrites  ore.  When 
the  fire  was  started  with  a  little  wood,  it  was  kept  up  by  the  heat  of  the 
burning  sulphur.  It  was  essential  that  no  rich  ore  was  included,  as,  on 
account  of  the  action  of  kernel  roasting,  lumps  of  rich  sulphide  of  copper 
were  formed  which  could  not  subsequently  be  dissolved.  Some  of  the 
roasted  ore  was  at  times  mixed  with  the  unroasted  ore,  for  the  extraction 
of  the  copper  with  the  ferric  chloride  solution.  The  reactions  between 
the  salt  and  ferrous  and  ferric  sulphates  in  the  heaps  may  be  expressed 
as  follows : 

2FeSO,-|-4NaCl  +  03=Fe203  +  2Na2S04-l-4Cl. 
Fe2(SO,)3-h6NaCH-03  =  Fe2034-3Na2SO,  +  6CL 

The  chlorine  liberated,  acting  on  the  iron  and  copper  sulphides,  pro- 
duces ferric  and  cupric  chlorides.  There  are  therefore  present,  FeSO^, 
Fe2(S04)3,  CUSO4,  FeClj,  and  CuClj,  with  an  excess  of  salt,  which  changes 
the  ferric  sulphate  into  the  ferric  chloride. 

An  interesting  modification  of  the  Doetsch  process  was  for  some  time 
carried  out  on  a  very  extensive  scale  at  Naya,  close  to  Rio  Tinto.  In 
this  method,  the  heaps,  made  in  the  ordinary  way,  as  soon  as  they  began 
to  give  off  sulphurous  fumes,  were  covered  up  with  a  fresh  quantity  of 
mineral,  partly  raw,  and  partly  roasted,  to  which  2  to  3  per  cent,  of  salt 
and  a  similar  proportion  of  manganese  dioxide  has  been  added.  The 
whole  was  formed  into  heaps  26  ft.  high  with  a  flat  top,  which  was  divided 
by  gutters  into  squares  of  25  ft.  The  remaining  operations  were  effected 
in  the  usual  way,  the  heaps  being  watered  at  intervals  for  months  and 
years,  the  copper  being  slowly  dissolved,  and  collected  at  the  bottom  of 
the  heaps.  It  was  necessary  to  break  up  the  surface  with  a  pick  at  inter- 
vals, to  prevent  it  from  becoming  impermeable  to  water.  The  sulphur- 
ous acid  gas  in  this  modification  of  the  process,  in  the  presence  of  steam 
formed  by  the  heat  developed  in  the  heaps,  produces  sulphuric  acid, 
which  acts  upon  the  oxides  in  the  crust  of  the  roasted  material.     The 


salt  and  manganese  dioxide  may,  jointly  with  the  sulphuric  acid,  evolve 
chlorine,  forming  ferric  chloride,  which  decomposes  the  sulphides  of 
copper  and  silver.  It  is  also  possible  that,  under  the  action  of  heat  and 
sulphuric  acid,  oxygen  is  evolved  which  acts  directly  upon  the  pyrites. 

The  precipitation  of  the  copper,  in  the  Doetsch  process,  was  effected 
in  a  series  of  tanks,  330  ft.  long  and  33  ft.  wide,  divided  into  10  parallel 
series,  receiving  a  uniform  supply  of  copper  solution.  The  total  length 
was  about  1300  ft.  with  a  difference  of  level  of  13  ft.,  which  gave  a 
sufficiently  rapid  flow.  Pig  iron,  as  run  from  the  furnace,  and  scrap  iron, 
were  used;  the  scrap  iron  was  put  into  baskets.  Every  10  days  the  iron 
was  removed,  and  after  scraping  to  collect  the  deposited  copper,  it  was 
returned  to  the  tanks.  The  consumption  of  iron  was  a  little  more  than 
one  ton  of  pig  iron  per  ton  of  copper  produced. 

The  precipitate,  as  collected,  was  very  impure,  containing  only 
65  per  cent,  to  70  per  cent,  copper,  the  remainder  being  ferric  oxide 
more  or  less  arsenical,  graphite  from  the  pig  iron,  silica,  etc.  After 
treatment  with  water  acidulated  with  sulphuric  acid  which  dissolved 
the  basic  ferric  arsenates  without  touching  the  copper,  it  was  passed 
over  4/10  in.  mesh  sieves  to  separate  pieces  of  cast  iron.  The  precipitate 
passed  through  the  sieves  was  washed  in  a  current  of  water,  where  it 
separated,  according  to  the  order  of  density,  into  "Cascara,"  or  copper; 
"Graphita,"  or  particles  of  coal  and  graphite;  and  "Pucha,"  a  fine  black 
sand.  The  copper  was  smelted  to  blister  copper;  the  graphita  was 
smelted  with  rich  ore,  and  the  Pucha  was  made  into  balls,  dried,  and 
also  smelted  with  the  ore. 

The  cost  of  producing  copper  by  the  Doetsch  process  has  been  esti- 
mated by  Cumenge,  according  to  the  results  which  he  accomplished  after 
a  campaign  of  four  months,  during  which  he  obtained  224  tons  of  cement 

Cost  of  leaoMng,  83 .  81 

Precipitation,  181 .  16 

General  expenses,  28 .  04 

Total  per  ton  of  cement  copper,  293 .  01 

Or  per  ton  of  pure  copper,  345 .  frances, 

or  $67.93 

The  extraction  at  the  expiration  of  four  months  amounted  to  1.34 
per  cent,  of  the  2.7  per  cent,  copper  contents.  1.12  ton  of  iron  was  re- 
quired to  precipitate  one  ton  of  cement  copper  which  is  equal  to  1.3  tons 
of  iron  to  one  ton  of  pure  copper. 

Dr.  0.  Froelich^  made  some  experiments  in  dissolving  copper  from 

'  Notes  sur  le  Rio  Tinto  by  M.  E.  Cumenge,  Annales  des  mines,  Vol.  XVCI. 
2  "  Metallurgie,"  1908,  p.  206. 



various  sulphide  ores  by  agitation  with  a  hot  solution  of  ferric  chloride, 
the  results  of  which  are  given  in  the  following  table: 

Characler  of 

Chalcocite  .  . 
Chalcocite. . . 

Chalcocite. . . 
Chalcocite. ,. 
Gray  copper. 

Extent  of 

Powder.  . 

19-32  in. 
13-32  in . 
5-32.  in 
Powder.  . 
Powder.  . 
Powder.  . 

Time  leached 

15   1/2 


66  1/2 

24  1/2 

Copper  cont. 

Copper  cont. 

of  ore, 

of  tails, 

per  cent. 

per  cent. 















Per  cent,  of 


In  these  experiments  the  chalcopyrite  was  first  subjected  to  a  tem- 
perature of  about  200°  C.  without  the  admission  of  air.  The  chalcocite 
was  treated  without  heating. 

The  Froelich  Process.' — In  this  process  the  ore  is  first  subjected 
in  the  absence  of  air  to  a  temperature  between  150  and  800°  C, 
whereby  the  loose  sulphur  is  driven  off.  The  pyrite  is  changed  by 
this  operation  in  its  chemical  composition  and  can  then  be  chlorinated 
much  quicker  and  better.  As  chlorinating  gases,  chlorine,  vapor  of 
hydrochloric  acid  and  of  ferric  chloride  are  used.  The  chlorine  and 
ferric  chloride  attack  the  sulphur  compounds  of  copper,  the  hydro- 
chloric acid  and  the  ferric  chloride  attack  the  ox'des  of  the  copper. 
The  proportion  of  the  gases  and  vapors  in  the  mixture  are  adjusted 
to  the  composition  of  the  copper  ore.  Steam  may  also  be  added 
to  the  chlorinating  mixture.  It  is  preferable  to  make  the  temper- 
ature of  the  chlorination  somewhat  higher  than  the  boiling  point  of  the 
ferric  chloride,  about  300°  C,  but  it  can  be  lower  if  the  copper  in  the  ore 
occurs  in  combination  with  sulphur.  In  this  case  only  chlorine  gas  is 
used.  If  iron  is  present  in  the  ore  it  is  chlorinated  together  with  the 
copper,  but  more  slowly,  and  the  process  is  facilitated  by  a  higher  tem- 
perature. In  order  to  regain  the  chlorine  combined  with  the  iron,  the 
ore  is  heated  to  about  300°  C.  or  more  after  chlorination,  and  a  certain 
quantity  of  air  is  introduced  during  this  operation. 

The  ferric  chloride  Is  then  evaporated,  and  by  the  air  separated  into 
oxide  of  iron  and  chlorine  gas,  both  of  which  are  collected.  The  chloride 
of  copper  is  not  changed  by  this  operation.  Then  the  ore  is  treated 
with  hot  water,  and  the  chloi'ide  of  copper  and,  perhaps,  a  residue  of 
ferric  chloride  are  extracted.  The  solution  is  then  introduced  into  a 
revolving  apparatus  containing  pieces  of  iron,  and  the  metallic  copper 
is  deposited  in  the  form  of  cement  copper.  The  solution  now  contains 
mainly  ferrous  chloride,  and  in  order  to  regain  the  chlorine  from  it  the 
'  U.  S.  Patent,  846,657,  March  12,  1907. 


solution  is  oxidized  in  a  rotating  drum  to  ferric  chloride  by  means 
of  an  air  blast.  Then  the  water,  and  the  wat^r  of  crystallization 
are  driven  off  by  heating,  and  by  increasing  the  temperature  over  the 
boiling  point  of  ferric  chloride  and  introducing  a  certain  quantity  of 
air,  chlorine  gas  and  ferrous  chloride  are  obtained. 

W.  L.  Austin^  suggests  the  following  method  for  treating  with  ferric 
chloride  pyritic  copper  ores,  containing  the  copper  largely  as  chalcocite. 
The  results  obtained  in  the  laboratory  have  been  very  satisfactory. 
The  novel  features  introduced  are  (1)  causing  the  lixiviant  to  rise  through 
the  ore  and  removing  in  an  apparatus  placed  outside  of  the  leaching  vat 
any  slimes  which  may  be  carried  over;  (2)  causing  the  lixiviant  to  cir- 
culate rapidly  through  the  material  treated  without  employing  moving 
parts  within  the  leaching  vat;  (3)  regeneration  of  the  lixiviant  by 
treating  it  with  chlorine  gas  produced  by  the  electrolysis  of  common 
salt  in  an  apparatus  specially  provided  for  that  purpose;  and  (4) 
cementation  with  the  aid  of  the  coke-iron  couple.  These  features  make 
it  possible  to  have  each  succeeding  stage  in  the  process  under  direct 
supervision,  and  avoids  complications  caused  by  attempting  to  carry  out 
two  distinct  operations  concurrently  in  the  same  apparatus.  It  also 
avoids  the  use  of  moving  parts  submerged  in  a  gritty  and  corrosive 

On  the  basis  of  a  plant  treating  100  tons  of  ore  in  24  hours,  and 
assuming  a  2  per  cent.'  ore  (40  lb.  copper  to  the  ton),  chlorine  at  $0.02 
per  pound,  arid  a  90  per  cent,  extraction,  the  following  estimated  cost  of 
producing  one  pound  of  refined  copper  by  this  method  is  derived: 

per  lb. 
Milling  operations,  comminution  of  the  ore  to  ten-mesh,    charging 
and  discharging  the  tanks,  elevating  liquors,  etc.,  total  $0.50 
per  ton  of  ore  treated,  $0 .  014 

Chlorine,  at  $0.02  per  lb.,  0.015 

Iron,  at  $30.02  per  ton  for  pig  delivered  at  works,  0.015 

Melting  the  precipitate  into  bars,  at  $8.00  per  ton  of  precipitate,         0 .  005 
Freight,  refining  charges,  and  selling  expenses,  0.014 

Repairs  and  renewals,  office  expenses,  etc.,  0.013 

Mining  operations — open  pit  work,  at  $0.50  per  ton  of  ore,  0.014 

Total  operating  expenses,  $0 .  090 

In  this  estimate  no  item  has  been  inserted  representing  interest  on 
cost  of  plant,  nor  amortization;  on  the  other  hand,  no  allowance  is  made 
for  precious  metals  recovered,  nor  for  possible  commercially  valuable 
'  Mines  and  Methods,  January  1911. 


bi-products.  The  figures  given  for  costs  of  reagents  consumed  are 
liberal  estimates. 

Ferrous  Chloride  Process. — The  action  of  ferrous  chloride  on  copper 
carbonate  was  demonstrated  by  experiments  of  Schaffner  and  Unger 
in  1862.  Some  years  later  Hunt  and  Douglas  based  a  process  on  the 
action  of  ferrous  chloride  on  copper  oxide  and  carbonate,  which  was 
for  some  time  working  on  a  commercial  scale  at  Ore  Knob,  North  Caro- 
lina, and  at  Phoenixville,  Pa. 

Cupric  oxide  and  cupric  carbonate  are  acted  upon  by  solutions  of 
ferrous  chloride,  forming  cupric  and  cuprous  chlorides  and  oxide  of  iron, 
and  in  the  case  of  carbonates,  there  is  also  set  free  carbon  dioxide, 
according  to  the  following  equations: 

3CuO  +  2Fe(  "I^  =  CuCl^  +  2(;uCl  +  FcO^. 
3CuC03  +  2FeCl2  =  ('uCl2  +  2CuCl+Fe203+3C(),. 

The  ferric  oxide  is  precipitated,  while  the  chlorides  of  copper  go 
into  solution.  Cuprous  chloride,  being  insoluble  in  water,  is  dissolved 
by  excess  of  other  metal  chlorides. 

If  the  ore  to  be  treated  is  a  sulphide,  or  a  matte,  the  copper  must 
first  be  converted  into  the  oxide  by  I'oasting.  Roasting  is  unnecessary 
with  oxidized  ores. 

The  ferrous  chloride  used  in  the  process,  may  be  produced  from 
common  salt  and  ferrous  sulphate,  which  by  double  decomposition  forms 
ferrous  chloride  and  sodium  sulphate.  After  the  solution  was  separated 
from  the  sodium  sulphate,  which  crystallized  out,  if  was  ready  for  use. 

When  copper  is  precipitated  from  a  cuprous  chloride  solution  by 
iron,  ferrous  chloride  is  formed;  when  both  cupric  and  cuprous  chlorides 
are  present,  ferric  chloride  is  first  formed, 

2CUCI2  +  2CuCl  +  2Fe  =  2FeCl3  +  4Cu. 

Ferric  chloride,  however,  reacts  with  iron  to  produce  ferrous  chloride 


The  silver  in  the  ore  is  converted  into  chloride  of  silver  by  the  cupric 
chloride,  and  is  dissolved  in  the  excess  of  other  metal  chlorides. 

As  some  of  the  copper  is  in  the  cuprous  condition,  relatively  less 
iron  is  required  to  precipitate  the  copper,  than  is  required  in  precipi- 
tating from  a  sulphate  solution.  The  objections  to  the  process  was  the 
formation  of  basic  salts,  and  the  difficulty  of  separating  the  solution 
from  the  residue,  on  account  of  the  precipitated  oxide  of  iron,  which 
clogged  the  filters.  Heat  is  not  necessary  in  dissolving  the  copper,  but 
it  hastens  the  process,  and  makes  the  extraction  more  thorough. 

The  silver,  in  solution  as  silver  chloride,  may  be  precipitated  witii 
the  copper,  or  separately  if  ilesired,  by  metallic  copper.  The  precipi- 
tation of  the  silver  by  metallic  copper  in  the  presence  of  chlorides  re- 



quires  that  the  whole  of  the  dissolved  copper  shoiild  be  in  the  cuprous 
condition.  The  clear  solution  containing  the  cupric  and  cuprous  chlo- 
rides and  the  silver  chloride  may  be  treated  with  sulphur  dioxide  to 
convert  the  cupric  chloride  into  the  cuprous  chloride,  and  with  liberation 
of  free  acid.  From  such  an  acid  solution  any  silver  present  is  readily 
and  completely  precipitated  by  metallic  copper,  after  which  the  whole 
of  the  dissolved  copper  may  be  precipitated  with  metallic  iron,  care 
being  taken  to  arrest  the  process  before  the  free  acid  begins  to  act  on 
the  iron.  In  this  way  a  solution  is  obtained  containing,  besides  the 
regenerated  ferrous  chloride,  a  considerable  amount  of  free  acid. 

When  copper  ores  containing  lime  are  being  treated,  there  is  the 
difficulty  to  contend  with  that  the  ferrous  chloride  and  calcium  carbonate, 
m  the  presence  of  air,  are  decomposed  into  calcium  chloride  and  ferric 

According  to  Hunt'  the  hydrous  silicate  of  copper  (chrysocoUa)  is, 
like  carbonate  of  copper,  completely  decomposed  by  a  hot  solution  of 
ferrous  chloride  with  common  salt. 

The  Ferrous  Chloride  Process  as  Carried  out  at  Ore  Knob,  Ashe  Co., 
N-  C.^ — At  Ore  Knob,  in  North  Carolina,  the  ore  was  crushed  to  40  mesh 
and  roasted.  The  average  composition  of  the  crude  ore,  as  taken  from 
the  two  shafts,  was  as  follows: 

No.  1 

No.  2 


11. 33  per  cent. 

13. 30 per  cent. 




Ferric  oxide, 















Carbonic  acid, 












Silioious  residue, 





Metallic  copper. 



This  ore  was  sorted  to  bring  the  average  copper  content  up  to  about 
12.0  per  cent.  The  average  composition  of  the  copper  in  the  roasted 
ore  is  represented  by  the  following  analysis : 

Copper  as  sulphate,  3 .  76  per  cent. 

Copper  as  oxide,  7 .  75  per  cent. 

Copper  as  sulphide,  0 .  39  per  cent. 

11. 90  per  cent. 
'  Trans.  A.  I.  M.  E.,  Vol.  X,  p.  12, 
^  T.  A.  I.  M.  E.,  Vol.  II,  p.  394,  E.  E.  Olcott. 

CHEMICAL  PROCESSES  .        227 

The  roasted  ore  was  conveyed  to  agitating  tanks,  which  were  eight 
in  number,  8  ft.  in  diameter,  and  5  ft.  deep,  with  raised  conical  bottoms. 
These  tanks  were  each  charged,  once  in  24  hours,  with  3000  lb.  of  roasted 
ore  and  1500  gallons  of  the  solution  of  ferrous  sulphate  and  common 
salt,  making  about  22°  B.,  and  heated  by  steam  to  160°  F.  This  mixture 
was  kept  in  agitation  for  eight  hours  by  means  of  suspended  stirrers, 
consisting  of  a  vertical  shaft  with  a  horizontal  blade  at  the  lower  extremity, 
while  at  the  top  was  attached  a  bevel  gear  which  gave  to  the  stirrers  a 
speed  of  25  revolutions  per  minute.  After  eight  hours  the  stirrers  were 
stopped,  and  the  contents  of  the  tank  allowed  to  settle  for  four  hours, 
when  the  clear  liquor  was  drawn  off  into  the  precipitating  tanks,  and  the 
remaining  portion  holding  in  suspension  ferric  oxide  and  particles  of 
gangue,  was  drawn  into  settling  tanks.  The  sands,  remaining,  were 
then  washed,  first  with  hot  strong  solution,  and  then  with  weaker  solu- 
tion. These  washing  liquors  were  allowed  to  settle  in  the  settling  tanks, 
when  the  clear  portion  was  drawn  off  into  precipitating  tanks  containing 
iron.  The  wet  sands  were  then  removed  from  the  agitators  to  leaching 
tanks  where  a  portion  of  the  adhering  solution,  containing  copper,  was 

The  slimes  were  allowed  to  accumulate  in  the  settling  tanks  till  they 
were  about  half  full,  when  they  were  washed  with  solution  and  water  till 
they  contained  about  1/2  of  1  per  cent,  of  copper.  They  were  then 
washed.  The  settling  tanks  were  20  in  number,  10  ft.  in  diameter  and 
5  ft.  deep. 

The  strong  liquors  from  the  agitators  were  capable  of  holding  50  lb. 
of  copper  in  solution  per  100  gallons,  but  weaker  solutions  were  desired 
as  they  lessened  the  risk  of  the  deposition  of  cuprous  chloride  by  cooling. 
30  lb.  of  copper  per  100  gallons  was  found  a  convenient  strength. 

The  hot  and  strongly  colored  liquors  were  run  through  launders  into 
the  precipitating  tanks.  These  were  12  in  number,  12  ft.  in  diameter 
and  5  ft.  deep,  containing  each,  12,000  lb.  of  scrap  iron.  The 
temperature  of  the  precipitating  tanks  was  maintained  at  160°  F.  by  the 
injection  of  steam.  From  12  to  18  hours  sufficed  to  precipitate  all  but  a 
trace  of  the  copper  from  the  liquors,  which  were  then  drawn  off  into  a 
lower  tank  and  from  there  pumped  into  the  stock  tanks  to  be  again 
used  on  a  fresh  portion  of  the  ore. 

For  the  precipitation  of  the  copper  wrought  iron  was  used.  The 
copper  was  removed  from  the  precipitating  tanks  when  they  contained 
from  4  to  5  tons  each.  The  consumption  of  iron  was  70  per  cent,  of  the 
pure  copper  produced.  The  cement  copper,  after  being  washed  and 
dried,  contained  generally  from  75  to  85  per  cent,  copper.  The  impuri- 
ties were  chiefly  ferric  oxide  and  earthy  gangue  matter. 

The  cement  copper  produced  at  Ore  Knob  costs  a  little  less  than 
8  cents  per  pound  of  copper,  including  all  expenses  for  mining,  treating 


and  packing.  Of  this  sum  nearly  two  cents  was  for  metallic  iron.  These 
costs  were  based  on  a  production  of  400,000  pounds  of  copper. 

Hunt  and  Douglas  Process. i— In  the  Hunt  and  Douglas  process  the 
copper  is  dissolved  as  sulphate  and  precipitated  as  the  insoluble  cuprous 
chloride.  The  cuprous  chloride  is  then  converted  into  metallic  copper  by 
replacement  with  iron. 

The  process  is  based  on  the  reaction  described  by  Wohler  between 
sulphur  dioxide  and  a  solution  of  cupric  chloride,  in  which  one  half  of 
the  chloride  is  eliminated  to  form  hydrochloric  acid,  and  with  the  simul- 
taneous formation  of  sulphuric  acid.     The  reaction  may  be  expressed: 

2CuCl2  +  SO2  +  2H2O  =  2CuCl  +  2HC1  +  H3SO  4. 

The  Hunt  and  Douglas  process  consists  of  the  following  essential 

1.  Roasting,  if  the  ore  is  a  sulphide. 

2. -Extracting  thecopper  from  the  oxidized  ore  with  dilute  sulphuric 
acid,  regenerated  in  the  operation  of  the  process. 

3.  Conversion  of  the  cupric  sulphate  into  cupric  chloride,  by  the 
addition  of  some  soluble  chloride,  such  as  sodium,  calcium  or  ferric 

4.  Conversion  of  the  soluble  cupric  chloride  into  the  insoluble  cuprous 
chloride  by  the  addition  of  sulphur  dioxide,  and  with  the  simultaneous 
regeneration  of  acid.  The  acid  being  used  in  the  second  step  to  dis- 
solve more  copper. 

5.  Conversion  of  the  precipitated  cuprous  chloride  into  cupric  oxide 
or  metallic  copper  on  the  addition  of  milk  of  lime  or  replacement  with 

In  practically  carrying  out  the  process,  the  matte,  or  ore  if  a  sulphide, 
is  roasted.  With  care  in  roasting  one  third  of  the  copper  should  be 
converted  into  the  sulphate,  which  is  soluble  in  water.  The  ore  is  then 
leached  with  dilute  sulphuric  acid,  regenerated  in  a  later  stage  in  the 
process.  The  copper  content  of  the  ore  is  therefore  dissolved  as  the 
cupric  sulphate,  CuSO^.  To  the  neutral  solution  of  cupric  sulphate  is 
then  added  enough  common  salt,  or  other  soluble  chloride,  to  convert  the 
copper  sulphate  in  the  solution  into  that  of  cupric  chloride.  The 
amount  of  copper  sulphate  being  determined,  salt  is  added  in  the  pro- 
portion of  58.5  parts  of  sodium  chloride  to  63.6  parts  of  copper  contained 
in  the  sulphate  solution.  The  salt  reacts  with  the  cupric  sulphate  to 
form  cupric  chloride  and  sodium  sulphate: 

CuSO,  +  2NaCl  =  CuCl,-hNa2S04, 
so  that  the  copper  in  the  solution,  after  the  application  of  the  sodium 
chloride  will  be  in  the  form  of  cupric  chloride,  containing  possibly  a 
little  cupric  sulphate. 

'  Trans.  A.  I.  M.  E.,  Vols.  X  and  XVI. 


Through  the  clear  hot  solution  of  cupric  chloride  is  then  driven  sul- 
phur dioxide,  derived  from  roasting  the  sulphide  ore.  The  sulphur 
dioxide  serves  to  convert  the  dissolved  copper  into  the  form  of  cuprous 
chloride,  with  the  liberation  of  the  amount  of  acid  which  was  previously 
combined  with  the  copper,  and  the  liberation  of  one  half  as  much  more 
acid  due  to  the  oxidation  of  the  absorbed  sulphur  dioxide. 

The  combined  reaction  between  the  copper  sulphate,  salt,  and  sul- 
phur dioxide  may  be  represented  by  the  equation : 

CuS0,-t-2NaCl-|-2SO2-h2H2O  =  2ru01-|-Na,.SO,-f-2H.,S(),. 

It  is  reasonably  certain,  however,  that  the  following  reaction  also  tiikes 
place,  whereby  a  certain  amount  of  hydrochloric  acid  is  also  produced: 

2CuCl2-|-S02-|-2H20  =  2CuCl-l-2HCl-t-H2SO,. 

The  cuprous  chloride  thus  obtained,  which  is  insoluble  in  water  or 
in  a  sulphate  solution,  quickly  settles  to  the  bottom  of  the  tank,  as  a 
white  crystalline  powder.  The  clear  acid  solution,  drawn  from  the  pre- 
cipitated cuprous  chloride,  is  again  applied  to  the  ore,  and  the  cycle 
repeated  as  often  as  necessary  to  get  the  desired  extraction. 

The  resulting  cuprous  chloride,  seperated  from  the  solution,  may 
then  be  precipitated  with  metallic  iron,  as  metalic  copper, 

2CuCl-hFe  =  2Cu-t-FeClj, 

or  with  milk  of  lime,  as  cuprous  oxide, 

2CuCl+Ca(OH)2  =  Cu20-t-CaCl2  +  H20, 

calcium  chloride  being  formed,  which  may  be  used  to  convert  the  sul- 
phate of  copper  into  the  chloride  instead  of  the  salt,  or  ferric  chloride. 

By  the  use  of  a  solvent  containing  only  a  small  portion  of  soluble 
chloride,  any  silver  in  the  ore  is  converted  into  the  chloride,  but  remains 
in  the  residue  and  may  be  extracted  therefrom  by  solution,  amalgama- 
tion, or  smelting. 

The  reaction  between  sulphur  dioxide  and  a  solution  of  cupric  chlo- 
ride goes  on  slowly  at  ordinary  temperatures',  but  is  very  rapid  between 
80  and  90°  C.  (176  to  194°  F.).  Solutions  of  sulphate  of  copper,  mixed 
with  an  equivalent  of  chloride  of  sodium,  and  holding  8  per  cent,  of 
copper,  after  being  treated  at  90  °  C.  with  an  excess  of  sulphur  dioxide, 
retain  less  than  1  per  cent,  of  the  dissolved  copper,  while  in  the  presence 
of  an  excess  of  sulphate  of  copper  and  sulphur  dioxide,  the  precipita- 
tion of  the  chlorine  from  chloride  of  sodium  is  nearly  complete. 

The  sulphur  dioxide,  from  the  roasting  furnace,  is  sufficiently  pure 
for  use.  A  Knowls  pump,  connected  for  the  purpose,  has  proved  an 
efficient  means  of  injecting  the  heated  gas  into  the  liquid. 

The  acid  liquors,  when  the  reaction  with  sulphurous  acid  is  com- 


pleted,  have  exchanged  their  bright  blue  color  for  a  pale  green,  and  now 
contain  in  solution  an  excess  of  sulphur  dioxide  which  must  be  got  rid  of 
before  using  it  to  dissolve  a  fresh  portion  of  copper.  This  may  be  effected 
by  keeping  back  a  small  portion  of  the  chlorinated  copper  solution,  and 
after  the  reduction  of  the  gas  is  complete,  as  may  be  shown  by  the  changed 
color  and  the  sulphurous  odor  of  the  liquid,  adding  the  reserve  portion 
thereto,  by  which  means  the  excess  of  siilphurous  acid  will  be  oxidized. 
The  larger  part  of  the  cuprous  chloride  separates  during  the  passage 
of  the  gas  but  a  furthur  portion  is  deposited  on  the  cooling  of  the 

The  excess  of  sulphurous  acid  may  also  be  got  rid  of  by  blowing  a 
current  of  hot  air  through  the  liquid  after  it  has  been  withdrawn  from 
the  precipitated  cuprous  chloride. 

Cuprous  chloride  is  quickly  transferred  into  cupric  oxychloride  by 
atmospheric  oxygen  and  when  dissolved  or  suspended  in  an  acid  liquid 
is  by  this  means  converted  into  a  cupric  salt,  which  may  be  again  reduced 
to  cuprous  chloride  by  the  action  of  sulphur  dioxide. 

If  the  ore  or  roasted  matte  contains  silver,  the  sulphate  of  copper, 
which  in  well-roasted  ore  should  be  about  one-third  of  the  copper  con- 
tent, is  first  dissolved  out  with  water,  taking  care,  however,  to  add  enough 
of  some  soluble  chloride  to  chlorinate  and  render  insoluble  any  sulphate 
of  silver  which  may  be  present.  From  the  clear  solution  thus  obtained, 
after  adding  the  requisite  amount  of  chloride  of  sodium,  the  copper  is 
precipitated  as  already  described,  by  the  action  of  sulphur  dioxide.  The 
resulting  acid  liquor,  freed  from  its  sulphur  dioxide,  is  now  used  to  dis- 
solve the  oxide  of  copper  in  the  ore,  the  process  being  aided  by  heat,  and 
if  the  formation  of  cuprous  chloride  is  feared  a  current  of  heated  air 
may  be  injected  and  made  the  means  of  agitating  the  mixture.  If 
the  ore  contains  the  silver  as  metal  or  oxidized  sulphide,  the  chloride  of 
copper  formed  is  a  good  agent  for  bringing  it  into  the  condition  of  silver 
chloride.  This  will  be  found  in  the  residue  after  the  extraction  of  the 
copper,  together  with  any  gold  which  may  be  present;  lead  as  sulphate, 
oxides  of  antimony  and  iron  and  earthy  matters.  Cobalt,  nickel  and 
zinc,  if  present,  will  however  be  dissolved,  and  not  being  precipitated 
by  sulphurous  acid,  will  by  successive  operations,  accumulate  in  the 
solutions  and  may  be  afterward  extracted.  From  the  residues,  the 
silver  may  readily  be  extracted  by  brine,  after  which  the  gold,  if  present, 
may  be  recovered  by  chlorination,  or  the  precious  metals  extracted 
together  from  the  residues  by  amalgamation. 

Chloride  of  silver  is  soluble  to  some  extent  in  a  solution  of  cupric 
chloride,  and  is  then  in  part  carried  down  with  the  cuprous  chloride  in 
the  precipitation  of  the  latter. 

The  cuprous  chloride  as  obtained  by  the  precipitation  with  sulphur 
dioxide  is  a  white  coarsely  crystalline  powder,  having  a  specific  gravity 


of  3.376  and  is  nearly  insoluble  in  water.  After  being  washed  from  the 
acid  liquid,  it  may  readily  be  reduced  by  placing  metallic  iron  in 
the  moist  cuprous  chloride,  which  should  be  covered  to  exclude  the  air. 
The  action  spreads  rapidly  through  the  precipitate,  so  that  a  single  mass 
of  iron,  within  a  few  hours,  will  change  a  considerable  volume  of  cuprous 
chloride,  around  it,  into  a  pure  spongy  metallic  copper.  Twice  the 
amount  of  copper  is,  theoretically,  precipitated  by  iron  from  a  cuprous 
than  from  a  cupric  solution.  45  lb.  of  iron  will  suffice  to  reduce  100  lb. 
of  copper  from  cuprous  chloride.  The  ferrous  chloride  which  remains  in 
the  solution  may  with  advantage  be  used  instead  of  sodium  chloride 
for  chlorinating  subsequent  solutions  of  copper  sulphate. 

Another  method  of  treating  the  cuprous  chloride,  consists  in  decom- 
posing it,  preferably  at  a  boiling  heat,  with  a  slight  excess  of  milk  of 
lime.  The  cuprous  chloride  is  by  this  means  converted  into  a  dense 
orange-red  cuprous  oxide,  which  after  being  washed  from  the  chloride  of 
calcium  in  a  filter  press  or  otherwise,  and  dried,  may  be  readily  reduced 
to  metallic  copper,  in  a  reverberatory  furnace.  For  this  reaction,  28.0 
parts  of  quick  lime  are  required  for  63.4  parts  of  copper,  and  the  result- 
ing chloride  of  calcium  may  be  used  instead  of  sodium  chloride  or  iron 
chloride  for  chlorinating  solutions  of  copper  sulphate.  In  this  case  there 
will  be  formed  an  insoluble  sulphate  of  lime,  while  the  free  sulphuric 
acid  of  the  solution  is  replaced  by  hydrochloric  acid. 

Later,  Douglas  proposed  electrolyzing  the  precipitated  solid  cuprous 
chloride,  to  deposit  metallic  copper,  and  use  the  chlorine  again  directly 
on  finely  crushed  matte.  A  description  of  this  will  be  found  under 
Electrolytic  Processes,  page  347. 

The  Hunt  and  Douglas  process  was  in  operation  for  many  years  on 
a  large  scale,  until  quite  recently,  at  Argentine,  Kansas,  on  copper  matte. 
The  description  here  given,  of  the  practical  operation  of  the  process,  at 
Argentine,  is  by  Ottakar  Hofmann. 

The  Hunt  and  Douglas  Process  at  Argentine,  Kansas.' — The  material 
treated  at  the  plant  of  The  Kansas  City  Smelting  and  Refining  Company, 
at  Argentine,  Kansas,  was  a  lead-copper  matte,  averaging: 

Copper,  39 .  55  per  cent. 

Lead,  12. 26 per  cent. 

Iron,  19  •  90  per  cent. 

Zinc,  1  •  88  per  cent. 

Manganese,  1  ■  01  per  cent. 

Sulphur,  21. 43  per  cent. 
Silver  varied  from  200  to  300  oz.  per  ton. 

This  matte  was  first  crushed  by  rock  breakers,  then  pulverized  in  a 
Krupp  ball  mill  to  pass  a  50-mesh  screen. 

The  roasting  was  done  in  two  Pearce  two-hearth  furnaces.     On  the 

'  Ottokar  Hofmann,  Mineral  Industry,  Vol.  XVII,  1908,  p.  296. 


upper  hearth  the  temperature  was  kept  as  low  as  the  heat  developed  by 
the  oxidation  of  the  sulphur  permitted.  No  fire  was  applied  except 
after  the  ore  had  passed  the  whole  circle  of  the  hearth  and  came  near  to 
the  slot  through  which  it  dropped  into  the  lower  hearth.  There  a  very 
gentle  fire  was  maintained  to  prevent  the  temperature  from  falling  too  low. 
The  best  results  were  obtained  by  regulating  the  roasting  on  the  upper 
hearth  so  that  the  material  commenced  to  ignite  when  it  had  moved  about 
8  ft.  from  the  point  at  which  it  had  entered  the  furnace.  By  observing  this 
precaution  the  roasting  was  so  much  advanced  by  the  time  the  material 
had  reached  the  drop-slot  that  the  oxidation  of  the  sulphur  did  not 
create  more  heat.  This  point  in  roasting  was  readily  observed  by 
stirring  the  charge;  if  the  particles  thrown  to  the  surface  brightened  and 
remained  so  for  a  short  while  the  oxidation  still  evolved  heat;  but  if 
these  particles  were  of  a  dead  red  color  and  began  to  darken  immediately, 
it  was  an  indication  that,  in  order  to  continue  the  oxidation,  heat  must 
be  applied.  It  was  found  to  be  of  the  greatest  importance  to  have  the 
roasting  well  advanced  when  the  material  left  the  upper  hearth.  When 
it  was  neglected  and  the  speed  of  the  feed  increased  so  that  the  matte, 
after  having  dropped  to  the  lower  hearth  still  created  heat  by  oxidation, 
the  finished  product  was  invariably  insufficiently  roasted.  It  was  endeav- 
ored to  maintain  a  gradually  increasing  temperature  in  the  lower  hearth  up 
to  the  point  of  discharge. 

In  order  to  regulate  the  final  heat,  tests  were  made  at  intervals  of 
the  material  before  and  after  it  passed  the  last  fire.  The  samples  were 
sifted  and  washed  in  a  small  dish  to  determine  if  any  cuprous  oxide  had 
been  formed.  The  presence  of  cuprous  oxide  is  readily  determined  by  its 
pink  color.  It  often  happened  that  although  the  material  was  free  from 
cuprous  oxide  before  passing  the  last  fire  place,  it  could  be  plainly  detected 
after  passing  it.  This  was  always  an  indication  that  the  fire  was  too 
hot.  It  was  important  to  avoid  this  condition  because  by  too  high  a 
temperature  cupric  sulphate,  of  which  quite  a  percentage  was  formed 
during  roasting,  was  decomposed  into  cuprous  oxide  and  sulphuric  acid, 
and  the  matte  was  discharged  before  the  cuprous  could  be  oxidized  to 
cupric  oxide.  This  test  had  to  be  made,  not  in  order  to  prevent  the  loss 
of  acid,  because  in  the  Hunt  and  Douglas  process  more  acid  is  made  than 
needed,  but  for  the  reason  that  when  cuprous  oxide  is  treated  with  dilute 
sulphuric  acid,  only  one-half  of  the  copper  can  be  dissolved  as  cupric  sul- 
phate; the  other  half  changes  into  metallic  copper,  which  being  insoluble, 
will  remain  in  the  residues. 

Even  with  the  greatest  care  it  was  impossible  to  roast  a  leady  matte 
free  from  small  lumps.  They  formed  in  the  very  early  period  of  the 
process  before  any  additional  heat  was  used,  but  as  a  rule,  being  usually 
porous,  they  were  found  well  roasted.  These  lumps,  however,  were  very 
undesirable  in  the  subsequent  operation,  as  they  retarded  the  solution 


of  the  cupric  oxide  in  dilute  sulphuric  acid.  The  roasted  matte,  therefore, 
had  to  be  crushed.  From  the  roasters  the  matte  was  automatically 
conveyed  to  the  revolving  cooling  tables,  then  fed  to  a  ball  mill  with 
50-mesh  screens.  From  the  ball  mill  it  was  elevated  and  convened  to 
the  storage  bins. 

Solution. — The  dissolving  of  the  cupric  oxide  had  to  be  done  in  agitat- 
ing tanks,  it  being  impracticable  to  conduct  the  operation  in  tanks  with 
filter  bottoms.  When  roasted  copper  matte  is  brought  in  contact  with 
dilute  sulphuric  acid,  or  even  water,  it  cements  and  hardens  to  such  an 
extent  that  it  cannot  be  handled  with  shovels  if  not  previously  loosened 
with  picks  or  bars.  This  hardening  of  the  material  prevents  to  a  great 
extent  the  free  percolation  of  the  solution;  this  causes  much  delay  and 
also  makes  the  discharging  of  the  tank  a  rather  difficult  task.  Even 
while  charging  the  agitating  tanks,  it  was  necessary  that  the  acid  solution 
be  kept  in  lively  motion  and  that  the  matte  be  introduced  in  a  gradual 
stream  and  not  charged  with  shovels;  otherwise  hard  chunks  were 

The  dissolving  was  done  in  wooden  agitating  stir  tanks,  12  ft.  in 
diameter  and  6  ft.  deep,  provided  with  a  strong  hard  wood  propeller 
which  entered  and  was  driven  from  above.  These  tanks  were  about 
two-thirds  filled  with  acid  solution,  containing  9  to  10  per  cent,  free  acid, 
to  which  some  wash  water  was  added.  This  acid  solution,  which  resulted 
in  the  process,  always  contained  2  to  2  1/2  per  cent,  copper.  The  agitator 
was  set  in  operation  and  a  jet  of  steam  introduced  through  a  lead  pipe 
entering  from  above  and  fastened  close  to  the  side  of  the  tank.  The 
roasted  matte  was  then  brought  from  the  storage  bins  in  cars,  which 
were  half  covered  and  provided  with  a  slot  through  which,  by  tilting  the 
cars,  the  matte  could  be  uniformly  charged  into  the  tank.  The  addition 
of  matte  to  the  dilute  sulphuric  acid  produces  considerable  heat,  which 
aids  the  solution  of  cupric  oxide  and  diminishes  the  amount  of  steam 
required  to  maintain  the  pulp  at  the  desired  temperature. 

After  a  certain  amount  of  matte  had  been  added,  the  pulp  was  fre- 
quently sampled;  these  samples  were  filtered  and  the  filtrate  tested  for 
free  acid.  When  the  solution  was  nearly  neutral  the  charging  of  the 
matte  was  stopped,  but  the  agitation  continued  until  the  solution  became 
neutral,  or  almost  neutral.  This  operation  was  performed  with  care  in 
order  to  avoid  an  excess  of  matte,  which  would  have  enriched  the  resi- 
dues with  copper.  In  mixing  the  acid  solution  with  wash  water,  care 
was  taken  to  have  enough  free  acid  present  so  that  the  resulting  solu- 
tion would  contain  from  6  per  cent,  to  7  per  cent,  copper. 

The  neutral  solution,  together  with  the  residues,  was  discharged 
through  an  outlet  near  the  bottom  of  the  stir  tank,  in  a  large  lead-lined 
cast  iron  pressure  tank;  thence  under  an  air  pressure  of  40  lb.  it  M'as 
forced   through  large  filter  presses  with   4x4  wooden  plates.     When 


charged,  each  press  was  capable  of  holding  5  tons  of  residues.  When  a 
press  was  filled,  compressed  air  was  applied  to  blow  out  as  much  as  pos- 
sible of  the  strong  solution  which  had  been  absorbed;  then  the  residues, 
while  still  in  the  press,  were  washed  with  water. 

Below  the  presses  there  were  two  rows  of  tanks,  one  to  receive  the 
strong  liquor,  the  other  the  wash  water.  Some  of  the  tanks  were  assigned 
to  the  stronger  portion  of  the  wash  water  which  went  back  to  the  process, 
while  the  remainder  were  used  as  collecting  and  settling  tanks  for  the 
weak  wash  water;  this  was  sent  to  scrap  iron  tanks  for  the  precipitation 
of  the  contained  copper.  The  strong  liquor  and  strong  wash  water 
tanks  were  connected  with  a  pressure  tank,  placed  on  a  lower  level,  by 
means  of  which  the  liquid  could  be  forced  to  the  stir  tank  level.  On 
opening  the  filter  presses  the  washed  residue  cakes  dropped  into  wooden 
push  cars;  they  were  then  wheeled  to  an  opening  in  the  press  floor,  through 
which  they  were  dumped  directly  into  railroad  cars.  The  residues, 
which  were  rich  in  silver  and  lead,  were  delivered  to  the  lead  smelting 

The  next  operation  was  to  chloridize  the  sulphate  solution.  For 
this  purpose  the  strong  solution  was  elevated,  by  means  of  a  pressure 
tank,  into  a  stir  tank  used  only  for  this  purpose.  The  solution  was  tested 
for  copper,  its  volume  measured,  the  total  copper  in  the  charge  calculated, 
and  as  much  common  salt  added  as  was  required  to  convert  the  copper 
present  into  cuprous  chloride  (58  parts  of  sodium  chloride  to  63.4  parts 
of  copper).  The  solution  was  agitated,  heated,  and  then  discharged 
into  storage  tanks;  from  these  tanks  the  chloridized  liquor  was  elevated 
and  charged  into  so-called  reducing  towers  for  the  treatment  with  sulphur 

Precipitation  of  the  Copper  with  Sulphur  Dioxide. — The  towers  in 
which  this  part  of  the  process  was  carried  out  were  made  of  steel  and 
lined  with  lead;  the  bottoms  were  cone-shaped.  There  were  four  towers. 
The  tops  were  tightly  closed,  provided  with  manholes,  inlet  pipes  for 
the  solution  and  an  outlet  pipe  for  the  gas.  The  cone-shaped  bottoms 
were  provided  with  gas  inlet  pipes,  a  steam  pipe  and  a  discharge  pipe. 
The  outlet  for  the  gas  was  connected  with  a  main  pipe  which  discharged 
into  a  wooden  stack.  The  gas  from  each  tower  passed  directly  into  the 
stack.  This  arrangement  caused  considerable  loss  of  gas.  By  tests  it 
was  found  that  a  gas  which  on  entering  a  tower  contained  7  per  cent, 
sulphur  dioxide,  contained  4  per  cent,  when  discharged,  so  that  there 
was  a  loss  of  57.1  per  cent.  Two  towers  were  then  connected  up  so  that 
the  gas,  after  passing  through  the  first,  was  made  to  pass  through  the 
next  tower.  By  this  alteration  the  loss  of  gas  was  reduced  to  25.4  per 
cent.,  equal  to  an  increased  utilization  of  31.7  per  cent.  This  experiment 
demonstrated  that  the  precipitation  of  the  copper  in  the  tower  under 
pressure  was  correspondingly  quicker  than  in  those  without  pressure. 


The  gas  was  furnished  by  three  revolving  cylindrical  furnaces,  of 
which  two  were  kept  in  operation  and  one  in  reserve.  These  furnaces 
were  lined  and  provided  with  ribs  for  continually  raising  the  ore  and 
dropping  it  in  a  shower.  The  front  end  of  the  furnace  was  closed,  but 
the  cover  was  provided  with  air  registers  and  two  discharge  openings 
for  the  roasted  ore;  the  latter  opened  and  closed  automatically  at  each 
revolution  of  the  furnace.  The  back  end  of  each  cylinder  projected  a 
few  inches  into  a  small  dust  chamber,  which  again  was  connected  with  a 
system  of  dust  chambers.  Through  the  roof  of  the  small  chamber,  in  a 
slanting  position,  entered  the  feed  pipe  of  the  furnace.  The  feeding  was 
done  by  a  very  short  screw  conveyor  which  could  be  regulated.  The 
material  consisted  of  iron  pyrites  concentrates,  rich  in  gold,  from 

The  dust  chamber  was  connected  with  a  heavy  lead  pipe  about  4  in. 
in  diameter  and  was  strengthened  with  iron  rings,  to  which  the  pipe  was 
fastened.  The  entire  length  of  this  pipe,  about  150  ft.,  was  cooled  by  a 
spray  of  water  so  that  the  gas  was  cooled  before  it  entered  the  pumps. 
There  were  two  double  acting  gas-pumps  of  which,  however,  one  was 
sufficient  to  do  the  work,  while  the  other  was  kept  in  reserve.  The 
cylinder  measured  27.5  in.  in  diameter  and  the  piston  had  a  stroke  of 
28  in.  so  that  each  stroke  furnished  about  19  cu.  ft.  of  gas.  The  speed 
had  to  be  regulated  so  as  to  get  a  good  roast  of  the  concentrates,  and  at 
the  same  time  produce  as  strong  a  gas  as  possible.  In  order  to  fulfill 
both  conditions,  it  was  found  that  the  resulting  gas  could  not  contain 
more  than  5  per  cent,  sulphur  dioxide.  Frequent  gas  tests  had  to  be 
made  in  order  to  maintain  this  percentage.  Sometimes  there  was  an 
increase  in  strength  up  to  7  per  cent.,  in  which  case  the  roasting  was  not 
satisfactory;  but  more  frequently  it  dropped  below  5  per  cent.,  which 
caused  a  slower  precipitation. 

The  pumps  forced  the  gas  through  a  lead  lined  receiver,  in  which 
a  great  deal  of  sulphuric  acid  condensed  and  had  to  be  drawn  off  daily. 
The  gas  entered  the  tower  under  a  heavy  perforated  lead  cone  which 
divided  it  into  small  bubbles.  Cuprous  chloride  was  precipitated  in 
white  crystals,  while  sulphuric  acid  was  set  free.  The  reaction  was  most 
energetic  in  the  beginning  while  the  solution  was  neutral  or  contained 
only  a  small  percentage  of  acid,  became  more  sluggish  in  proportion  as 
the  percentage  of  acid  increased,  and  stopped  entirely  when  the  copper 
contents  of  the  liquor  was  reduced  to  2  or  2  1/2  per  cent.  This  remain- 
ing copper  could  not  be  reduced  no  matter  how  long  the  charge  was  kept 
under  treatment  with  the  gas.  The  acid  continued  to  increase  slowly, 
but  the  copper  did  not  diminish.  It  is  possible  that  the  formation  of 
hydrochloric  acid  accounts  for  the  copper  not  being  precipitated.  The 
more  hydrochloric  acid  there  is  formed,  the  more  copper  will  remain  in 
solution.     The  hydrochloric  acid  dissolves  cuprous  chloride. 


Sodium  chloride  as  a  chloridizer  for  the  sulphate  solution  is  not  the 
most  suitable  chemical  for  the  process,  as  a  large  quantity  of  sodium 
sulphate  is  formed  which  goes  into  solution.  As  it  is  necessary  to  use  the 
solution  over  and  over  again,  on  account  of  the  sulphuric  acid  which  is 
formed  therein,  more  and  more  sodium  sulphate  is  formed.  In  a  short 
time  the  solution  becomes  saturated  with  this  salt  so  that  it- crystallizes 
out  whenever  conditions  are  favorable.  This  happens  at  different 
stages  of  the  process,  causing  much  annoyance  and  lessening  the  merit  of 
the  process.  It  was  especially  aggravating  in  the  operation  of  the  filter 
presses.  When  the  matte  residues,  together  with  the  strong  liquor, 
were  forced  into  the  presses,  the  filtration  in  the  beginning  was  free  and 
satisfactory,  but  soon  grew  less  so  until  it  finally,  stopped  entirely, 
although  the  press  was  not  one-quarter  filled  with  residues.  On  opening 
the  press  it  was  found  that  the  chambers  were  partly  filled  with  a  sloppy 
mass  containing  many  fine  crystals,  while  the  filter  cloth  was  densely 
covered  with  them.  The  only  way  of  cleaning  the  press  was  to  force 
water  through  it.  This  not  only  caused  much  delay,  especially  as  this 
application  of  water  had  to  be  repeated,  but  caused  the  making  of  a 
large  quantity  of  wash  water,  from  which  the  copper  had  to  be  pre- 
cipitated with  scrap  iron.  Sometimes  it  happened  that  the  press  could 
be  filled  without  any  trouble,  in  fact  the  chamber  filling  was  quite  firm; 
but  as  soon  as  water  was  used  to  wash  the  residues  the  filling  shrunk 
in  volume,  and  the  frames  which  previouslj^  were  quite  full,  after  washing 
were  only  a  little  over  half  filled. 

Calcium  chloride,  formed  in  converting  the  cuprous  chloride  into 
cuprous  oxide  by  boiling  with  milk  of  lime,  was  then  tried.  The  result- 
ing calcium  chloride  solution,  however,  was  not  strong  enough  to  be  used 
directly,  containing  only  9  per  cent,  chlorine;  it  was  therefore  concen- 
trated. In  using  calcium  chloride,  cupric  chloride  and  calcium  sulphate 
were  formed,  the  latter  being  precipitated.  Though  this  chloridizer 
made  necessary  an  additional  filter  press  operation,  to  separate  the 
calcium  sulphate  from  the  solution,  it  was  by  far  preferable  to  salt,  as 
it  left  a  clean  solution,  free  from  undesirable  salts.  It  was  found,  how- 
ever, that  a  sulphate  solution,  chloridized  with  calcium  chloride,  only 
about  half  of  the  copper  in  solution  could  be  precipitated  as  cuprous 
chloride  with  sulphur  dioxide.  "When  the  change  from  salt  to  calcium 
chloride  was  made,  a  new  acid  solution,  free  from  sodium  sulphate,  was 
used.  Several  attempts  were  made  with  the  same  unsatisfactory 
results.  It  was  finally  decided  to  add  salt  to  the  new  solution  and  the 
results  were  at  once  better.  From  that 'time  on  chloridizing  was  done  so 
that  three-fourths  of  the  required  chlorine  was  derived  from  calcium 
chloride  and  one-fourth  from  sodium  chloride.  After  adopting  this 
proportion  there  was  but  little  trouble  with  the  presses;  the  filtration 
was  free. 


The  calcium  chloride  solution  obtained  in  converting  the  cuprous 
chloride  into  cuprous  oxide  by  boiling  with  milk  of  lime,  was  concen- 
trated liy  preparing  milk  of  lime  with  water,  allowing  the  lime  to  settle, 
decanting  the  clear  water  and  replacing  the  same  by  weak  calcium 
chloride  solution.  The  condition  of  the  slacked  lime  was  not  changed 
and  by  repeating  this  procedure  the  proportion  of  calcium  chloride  in 
the  solution  was  increased  from  9  per  cent,  to  24  per  cent. 

When  the  precipitation  in  the  towers  was  completed,  the  cuprous 
chloride,  together  with  the  acid  solution,  was  discharged  into  a  system  of 
seven  cone-shaped  lead-lined  iron  tanks.  These  tanks  were  so  arranged 
that  the  liquor  flowed  from  one  to  the  other,  and  from  the  last  into 
special  charge  tanks,  to  be  in  readiness  to  dissolve  a  fresh  lot  of  matte. 
These  cone-shaped  tanks  served  a  double  purpose;  to  give  the  cuprous 
chloride  an  opportunity  to  settle,  and  to  cool  the  solution.  This  liquor 
when  hot  holds  in  solution  a  large  amount  of  cuprous  chloride  which 
precipitates  out  as  the  temperature  falls.  The  temperature  of  the  liquor 
in  the  first  cone  was  50. ,5°  C,  and  in  the  six  following  cones  it  was,  48.0, 
47.0,  45.0,  39.5,  and  37.0°  C,  so  that  the  temperature  of  the  last  cone 
was  19.5°  lower  than  that  of  the  first.  The  cooling  proved  to  be  suffi 
cient,  as  no  further  precipitation  took  place  in  the.  storage  tanks. 

Below  the  level  of  the  cones  were  three  vacuum  filters,  into  which  the 
former  (iould  be  discharged  by  opening  the  bottom  valve.  Into  these 
filters  the  cuprous  chloride  was  allowed  to  drain,  then  receiving  a  thorough 
washing.  The  washed  precipitate  was  then  converted  into  cuprous 

Conversion  of  the  Cuprous  Chloride  into  Cuprous  Oxide. — The  con- 
version was  done  with  milk  of  lime.  The  lime  was  slacked  in  a  flat  box 
and  collected  in  settling  tanks.  From  these  tanks  the  milk  of  lime  of 
proper  consistency  was  charged,  by  means  of  a  steam  syphon,  into  a  stir 
tank  and  heated  with  a  jet  of  steam.  The  washed  cuprous  chloride  was 
gradually  charged;  its  color  changed  from  white  to  red.  Cuprous  oxide 
and  calcium  chloride  were  formed.  As  calcium  chloride  dissolves 
cuprous  chloride,  the  calcium  chloride  which  was  formed  in  proportion 
as  the  conversion  progresses  will  dissolve  some  of  the  freshly  charged 
cuprous  chloride.  On  this  property  of  calcium  chloride  was  based  the 
test  by  which  the  conversion  is  conducted.  After  a  certain  amount  of 
cuprous  chloride  had  been  added  to  the  milk  of  lime,  charging  was  inter- 
rupted, the  stirrer,  however,  being  kept  in  motion.  About  10  minutes 
later,  a  sample  was  taken  in  a  wide  necked  bottle  suspended  by  a  copper 
wire.  Part  of  this  sample  was  filtered  and  nitric  acid  and  then  ammonia 
added  to  the  filtrate.  If  the  blue  color  appeai-ed,  some  cuprous  chloride 
was  still  dissolved  in  the  calcium  chloride  solution.  The  agitation  was 
then  continued  for  15  or  20  minutes,  when  another  sample  was  taken. 
If  the  blue  color  appeared  again,  more  milk  of  lime  was  gradually  added 


and  at  intervals.  After  each  interval  a  test* was  made  until  the  blue 
color  ceased  to  appear.  The  last  part  of  the  operation  had  to  be  con- 
ducted carefully  to  avoid  an  excess  of  lime. 

The  pulp  consisting  of  cuprous  oxide  and  calcium  chloride  was 
forced,  by  means  of  a  double  acting  pump,  into  a  Johnson  iron  filter 
press,  where  it  was  washed.  The  cakes  were  dumped  on  a  lower  floor 
and  dried  on  steam  slabs.  When  dry,  the  cuprous  oxide  was  carted  to 
the  copper  smelting  department  and  dumped  into  bins  conveniently 
arranged  on  the  charge  floor  of  a  cupola  furnace.  In  this  furnace  it  was 
reduced  to  metallic  copper. 

The  cuprous  oxide  always  contained  from  4  to  5,  and  sometimes  as 
much  as  11  oz.,  silver  per  ton.  Some  of  the  silver  undoubtedly  came  in 
with  fine  particles  of  matte  residues,  which  were  still  suspended  in  the 
solution  when  it  was  charged  into  the  towers  for  treatment  with  sulphur 
dioxide,  although  the  resulting  cuprous  chloride  was  clear  white  and 
did  not  show  any  coloration.  However,  after  more  settling  tanks  were 
inserted  for  the  solution,  the  cuprous  oxide  contained  considerably  less 
and  more  uniform  amounts  of  silver.  A  sample  of  cuprous  oxide  con- 
taining 6.25  oz.  silver  per  ton,  when  leached  with  a  solution  of  sodium 
hyposulphite,  still  contained  5.5  oz.  silver  per  ton,  so  that  only  0.75  oz. 
per  ton  could  be  extracted  by  that  solution.  This  test  was  made  with 
cuprous  oxide  produced  before  the  additional  settling  tanks  were  in  use; 
afterward  the  cuprous  oxide  did  not  contain  over  2  to  3  oz.  silver  per  ton. 
The  resulting  calcium  chloride  solution  gave  with  sodium  sulphide  a 
dark  precipitate  which  consisted  mostly  of  lead  sulphide,  with  only  a 
trace  of  copper  and  no  silver. 

In  smelting  the  cuprous  oxide  in  the  cupola,  very  strong  and  offen- 
sive fumes  were  formed.  These  fumes  were  white,  but  when  very  strong 
assumed  a  reddish  tinge.  They  consisted  chiefly  of  volatilized  cuprous 
and  cupric  chloride,  some  hydrochloric  acid  and  flue  dust  of  cuprous 
oxide.  An  investigation  showed  that  the  cuprous  oxide  still  contained 
1  to  2  per  cent,  chlorine,  notwithstanding  the  fact  that  it  was  subjected 
to  a  very  thorough  washing  in  the  press.  This  chlorine  could  not  be 
removed  or  reduced  even  by  an  extended  washing.  This  was  not  due  to 
the  presence  of  cuprous  chloride,  caused  by  an  insufficient  quantity  of 
lime  being  used  in  the  conversion,  for,  even  if  for  the  sake  of  information 
an  excess  of  lime  was  used  and  an  unusually  long  time  given  for  conver- 
sion, the  above  stated  percentage  of  chlorine  was  always  found  in  the 
cuprous  oxide. 

The  obnoxious  character  of  the  furnace  gases  was  destroyed  by 
passing  them  through  a  shower  of  milk  of  lime.  A  tower  was  arranged 
which  was  provided  at  different  levels  with  strong  wooden  grates.  At 
the  foot  of  this  tower  tanks  were  arranged  for  making  and  receiving  milk 
of  lime.     Two  of  these  receiving  tanks  were  connected  with  a  force 


pump.  The  flue  was  connected  with  the  wooden  tower.  Coarse  lime 
rock  was  placed  on  the  different  grates  to  detain  the  milk  of  lime  in  its 
downward  course  as  long  as  possible.  At  the  top  several  perforated 
pipes,  through  which  the  milk  of  lime  was  forced  by  means  of  the  pumps, 
were  so  arranged  as  to  furnish  an  even  spray.  The  bottom  of  the  tower 
was  made  tight  and  the  outlet  made  to  convey  the  stream  into  one  of  the 
other  tanks,  so  that  the  milk  of  lime  could  be  passed  through  the  tower 
as  often  as  desired. 

The  effect  of  the  milk  of  lime  was  very  gratifying.  The  strong 
offensive  odor  of  the  gases  disappeared  entirely.  The  color  of  the  milk 
of  lime  turned  gradually  darker  and  became  finally  olive  green  and  very 
rich  in  copper.  No  copper  escaped  with  the  gases.  When  the  above 
green  pulp  was  filtered  the  filtrate  contained  11/2  per  cent,  to  2  per  cent, 
chlorine.  The  evaporation  was  great  and  water  had  to  be  added  fre- 
quently to  maintain  the  same  volume  of  precipitate.  This  method 
proved  itself  successful  and  was  finally  permanently  installed. 

Treatment  of  the  Wash  Water. — In  the  course  of  the  process  a  great 
deal  of  wash  water  was  made,  principally  from  washing  the  copper  matte 
residues  and  the  cuprous  chloride.  The  latter,  which  contained  from 
1  to  2  per  cent,  copper,  was  collected  in  a  number  of  large  tanks,  from 
which  it  was  drawn  to  be  subjected  to  special  treatment  for  recovering 
the  copper.  Wash  water  containing  2  per  cent,  copper  and  as  much  as 
practicable  of  the  weaker  portion,  went  back  to  the  process  and  was 
used  instead  of  water. 

To  produce  a  clean  cement  copper  free  from  chlorine,  a  trough  was 
constructed,  in  sections,  abous  200  ft.  long  and  12  in.  wide  and  14  in. 
deep.  All  sections  were  placed  horizontally,  but  each  succeeding  one 
was  placed  three  inches  lower.  The  outlet  of  each  section  was  two  inches 
lower  than  the  inlet.  In  some  of  the  sections  the  compartment  was  made 
by  inserting  across  the  width  of  the  trough  two  boards  6  inches  wide. 
These  were  placed  about  12  in.  apart  to  allow  for  the  insertion  of  a  steam 
jet.  This  part  of  the  trough  was  tightly  covered  for  2  ft.  on  either  side 
of  the  jet  to  prevent  the  solution  from  being  splashed  out  by  the  steam. 
The  200  ft.  of  trough  was  arranged  in  U  shape  to  avoid  too  long  a  build- 
ing, and  to  make  the  handling  of  the  material  easier.  After  passing 
through  the  last  section,  the  solution  flowed  through  a  few  scrap  iron 
tanks  to  precipitate  any  copper  which  might  be  present  as  sulphate. 
Each  section  of  the  long  trough  was  charged  about  4  in.  deep  with 
cement  copper,  evenly  spread.  The  wash  water  from  the  main  depart- 
ment flowed  into  a  circular  tank  8  ft.  in  diameter  and  5  ft.  deep,  thence 
through  an  overflow  into  the  first  section  of  the  long  trough.  The  pur- 
pose of  the  circular  tank  was  to  heat  the  wash  water  by  means  of  steam 
jets  before  it  entered  the  long  trough.  It  gave  an  additional,  though 
not  very  effective,  opportunity  for  the  settling  of  particles  of  matte 


residues  which  might  not  have  settled  in  the  proper  wash  water  storage 
and  settling  tanks.  As  the  wash  water  resulting  from  the  different  oper- 
ations of  the  process  contained  cupric  sulphate  in  addition  to  cupric 
chloride,  salt  water  was  added  to  the  storage  tanks  to  convert  the  cupric 
sulphate  into  cupric  chloride. 

By  passing  the  wash  water  through  and  over  the  cement  copper  the 
copper  was  precipitated  as  cuprous  chloride.  Once  a  day  the  cement 
copper  in  the  troughs  was  gently  worked  in  order  to  bring  fresh  particles 
to  the  surface.  At  one  side  and  below  the  level  of  the  troughs,  a  re- 
volving barrel,  10  ft.  long  and  6  ft.  in  diameter  was  erected;  when  a  large 
part  of  the  cement  copper  in  the  troughs  had  changed  to  cuprous  chlo- 
ride, it  was  removed  from  the  troughs  and  charged  into  the  barrels,  to 
which  small  scrap  iron,  water  and  salt  were  added.  By  means  of  a  steam 
pipe,  the  pulp  was  slightly  heated  to  start  the  reaction;  then  the  steam 
was  turned  off.  This  pipe  entered  the  barrel  through  one  of  its  axles 
and  was  bent  downward  to  reach  into  the  pulp;  it  was  kept  in  position 
by  a  stuffing  box.  Very  soon  an  energetic  reaction  took  place;  the  heat 
developed  causing  the  pulp  to  boil  violently.  The  steam  found  an  out- 
let through  another  pipe  inserted  through  the  opposite  axle.  The  salt 
was  added  to  dissolve  some  of  the  cuprous  chloride;  this  caused  the 
reaction  to  start  more  quickly.  The  cuprous  chloride  changed  to  cement 
copper  and  the  iron  into  ferrous  chloride.  The  latter  had  the  same 
effect  as  salt  and  dissolved  cuprous  chloride,  thus  assisting  the  process. 

After  the  steam  ceased  to  escape,  the  barrel  was  stopped  and  a 
sample  taken.  The  sample  was  filtered  and  the  filtrate  tested.  When 
the  blue  color  could  no  longer  be  obtained  with  nitric  acid  and  ammonia, 
the  pulp  was  ready  to  be  discharged.  The  conversion  was  completed 
in  from  6  to  10  hours.  Below  the  barrel  was  a  square  tank,  with  a 
filter  bottom,  fitted  on  two  opposite  sides  with  rails  which  extended 
beyond  the  tank.  On  this  track  was  an  8-mesh  screen  fastened  to  a 
wooden  frame  and  provided  with  four  wheels.  The  screen  covered  the 
whole  top  of  the  tank.  In  discharging  the  barrel,  the  copper  cover  of 
the  manhole  was  removed  and  the  barrel  gradually  turned  by  means 
of  a  crow-bar.  When  the  charge  was  out,  the  inside  of  the  barrel  was 
rinsed  with  water.  The  screen  retained  pieces  of  iron,  while  the  cement 
copper  and  solution  passed  through  to  the  filter  tank.  By  means  of  a 
stream  of  water,  the  iron  on  the  screen  was  separated  from  adhering 
cement  copper  and  returned  to  the  barrel  to  serve  as  part  of  the  next 
charge.  The  screen  on  wheels  was  pushed  away  from  the  tank,  the 
outlet  under  the  filter  opened,  and  the  iron  solution  allowed  to  drain  off. 
Then  warm  dilute  sulphuric  acid  was  permitted  to  flow  in.  As  soon  as 
the  acid  appeared  at  the  outlet  the  latter  was  closed.  The  acid  was 
applied  to  remove  basic  salts,  to  prevent  their  formation  as  far  as  pos- 
sible, and  to  dissolve  any  small  pieces  of  iron  which  had  passed  through 


the  screen.  After  several  hours  the  acid  was  removed  to  a  special  tank 
to  be  used  again.  The  copper  was  then  thoroughly  washed  in  the  tank 
to  free  it  from  acid.  The  resulting  cement  copper  was  of  a  very  clear  color 
and  unusually  pure,  containing  99  per  cent,  copper  and  but  a  trace  of 
arsenic.     It  was  melted  in  the  refining  furnace. 

This  method  gave  such  good  results  that  three  more  barrels  were 
erected.  The  method  eliminated  the  conversion  of  cuprous  choloride 
into  cuprous  oxide  by  the  milk  of  lime,  the  smelting  of  the  product  in 
the  cupola  furnace,  the  treatment  of  the  furnace  gases,  the  very  unclean 
manipulation  of  the  scrap  iron  tanks,  and  the  additional  treatment  of 
the  cement  copper  to  free  it  from  cuprous  chloride. 

Modification  of  the  Hunt  and  Douglas  Process. — To  simplify  the  op- 
erations and  to  avoid  the  saturation  of  the  solution  with  sodium  sul- 
phate and  its  attendant  disadvantages,  Hofmann  worked  out  and 
successfully  introduced  the  following  modus  operandi: 

1.  The  process  was  started  with  a  stock  of  dilute  sulphuric  acid. 
By  treating  the  roasted  matte  in  the  usual  way  in  the  stir  tanks,  a  clean 
sulphate  solution  was  obtained  which  filtered  well  in  the  presses. 

2.  The  sulphate  solution  was  chloridized  with  hydrochloric  acid,  of 
which  in  starting,  a  stock  on  hand  was  required.  By  chloridizing  the 
sulphate  solution  with  hydrochloric  acid,  cupric  chloride  is  formed  and 
sulphuric  acid  set  free.  No  foreign  salts  are  introduced  and  the  solution 
remains  clean,  while  by  chloridizing  with  sodium  chloride  the  solution 
becomes  quickly  saturated  with  sodium  sulphate. 

3.  The  cupric  chloride  solution  containing  the  liberated  sulphuric 
acid  was  then  treated  in  a  stir  tank  with  cement  copper.  The  cupric 
chloride  by  the  action  of  metallic  copper  is  reduced  to  cuprous  chloride, 
while  the  sulphuric  acid  remains  unchanged.  A  steam  jet  is  used  to 
hasten  the  reaction.  An  excess  of  cement  copper  serves  the  same  pur- 
pose. When  the  filtrate  of  a  sample  shows  no  reaction  for  copper,  the 
operation  is  completed  and  the  pulp  is  drawn  into  a  pressure  tank  and 
forced  through  a  filter  press.  The  filtrate,  which  is  now  a  clean  sulphuric 
acid  solution,  is  elevated  to  storagfe  tanks,  whence  it  is  used  as  required 
to  dissolve  the  cupric  oxide  of  a  new  lot  of  roasted  matte.  To  produce 
as  little  wash  water  as  possible,  the  solution  absorbed  by  the  matte 
residues  and  by  the  cuprous  chloride  is  forced  out  by  compressed  air;  this 
works  very  well,  as  the  filtering  capacity  of  both  materials  is  not 
lessened  by  the  formation  of  crystals.  For  the  same  reason  the  subsequent 
Avashing  is  quickly  done,  requiring  comparatively  but  very  little  water 
to  accomplish  it. 

4.  The  washed  cuprous  chloride  was  treated  in  revolving  barrels 
in  the  same  manner  as  described  above,  but  care  was  taken  that  no  more 
water  was  added  than  necessary,  so  that  as  strong  a  ferrous  chloride 
solution  was  produced  as  practicable. 



5.  The  ferrous  chloride  solution  was  evaporated  in  the  same  iron 
pans  which  were  formerly  used  for  concentrating  the  calcium  chloride 

6.  The  solid  ferrous  chloride  was  charged  into  retorts,  which  were 
provided  with  water  for  steam  and  air.  When  heated  heavy  fximes  of 
hydrochloric  acid  were  formed;  these  were  passed  through  a  cooling 
arrangement,  in  which  a  large  portion  was  condensed.  This  condensed 
acid  was  strong  and  contained  35.6  per  cent,  chlorine.  The  acid  fumes 
which  were  not  condensed  were  made  to  pass  through  two  towers  made 
of  stoneware  pipes  and  filled  with  coke.  The  gas  escaping  from  the 
first  tower  entered  at  the  bottom  of  the  second.  To  avoid  the  accu- 
mulation of  too  much  water  in  the  stock  solution,  cupric  sulphate  solu- 
tion was  used  instead  of  water  as  a  spray  for  the  coke,  thus  chloridizing 
the  solution.  The  solution,  as  a  rule,  after  passing  through  the  towers, 
contained  an  excess  of  hydrochloric  acid.  This  condition,  however, 
was  properly  adjusted  by  adding  sulphate  solution  before  the  treatment 
with  cement  copper. 

7.  The  resulting  cement  copper  was  very  pure,  containing  from 
90  to  94  per  cent,  and  more  copper.  This  was  smelted  in  a  refining 
furnace;  no  obnoxious  fumes  were  evolved  as  in  the  case  of  cuprous 
chloride.  There  was,  of  course,  a  loss  of  sulphuric  acid  as  well  as  of 
hydrochloric  acid  which  was  caused  chiefly  by  the  wash  water;  these  had 
to  be  replaced  from  time  to  time  to  keep  up  the  volume  of  stock  solution. 
However,  this  shortage  was  not  great  and  the  loss  of  replacing  it  was 
far  less  than  that  of  the  eliminated  operations. 

This  modified  process  was  used  for  some  time  until  Hofmann 
received  instructions  from  the  company  for  the  necessary  alterations  of 
the  works  to  prepare  for  the  more  profitable  manufacture  of  blue  vitriol. 
This  was  done  by  Hofmann's  method  of  producing  this  material 
direct  from  the  roasted  copper  matte,  described  in  chapter  18. 

Copper  Extraction  at  Falun,  Sweden. ' — The  old  copper  mine  at  Falun 
produces  two  classes  of  minerals  known  as  hard  and  soft  pyrites  respec- 
tively. The  former,  consisting  of  mixtures  of  quartz  and  copper  pyrites, 
contains  about  3  1/2  per  cent,  and  the  latter,  which  is  mainly  composed 
of  iron  pyrites,  about  1  per  cent,  of  copper.  The  hard  ore  is  roasted  in 
heaps,  about  10  per  cent,  of  the  sulphur  being  driven  off,  while  the  soft 
pyrites  is  treated  in  sulphuric  acid  works,  about  30  per  cent .  of  the 
sulphur  being  utilized. 

The  burnt  residues,  to  the  extent  of  about  45  tons  hard,  and  12  tons 
soft,  per  day,  are  mixed  with  14  and  10  per  cent,  of  salt  respectively, 
and  ground  in  ball  mills.  The  hard  ore  mixture  is  then  subjected  to 
a  chloridizing  roasting  in  a  White-Howell  roasting  furnace,  in  which 

^E.  and  M.  J.,  Aug.  30,  1902;  Berg  and  Huttenmanisohe  Zeitung,  1902. 


15  tons  are  roasted  in  24  hours.  The  soft  ore  is  roasted  in  a  double  bed 
roaster  in  which  7  tons  are  worked  through  daily. 

When  complete  chloridization  has  been  effected,  the  roasted  material 
is  transferred  to  wooden  vats,  where  the  cupric  chloride  is  dissolved  out 
by  weak  sulphuric  or  hydrochloric  acid,  the  latter  being  obtained  by 
condensing  the  waste  gases  from  the  roasters. 

After  the  ore  has  been  treated  with  the  dilute  acid  solution,  it  is 
washed  with  clear  water.  The  wash  water  is  pumped  back  and  used  as 
a  second  liquor  on  a  following  charge. 

The  solution  from  the  ore  contains  all  the  copper,  bismuth,  selenium 
and  silver,  together  with  a  portion  of  the  gold  contained  in  the  ore. 
The  remainder  of  the  gold  is  extracted  later  by  washing  the  residues 
with  chlorine  water. 

The  copper  solution  is  precipitated  by  scrap  iron  as  cement  copper, 
which,  together  with  the  associated  mud  and  iron  salts,  is  smelted  and 
granulated  for  conversion  into  copper  sulphate  by  means  of  dilute  sul- 
phuric acid  and  air  in  the  ordinary  way. 

The  residual  mud  from  the  copper  sulphate  crystallizers,  containing 
gold,  silver,  selenium,  and  bismuth  is  dried  and  smelted  with  litharge, 
soda  and  sawdust  to  collect  the  precious  metals  into  lead.  The  gold- 
bearing  solution  from  the  chlorine  extraction  is  reduced  by  adding  a 
portion  of  the  original  copper  extracting  solution  containing  ferrous  chlo- 
ride, which  reduces  the  chloride  of  gold,  producing  metallic  gold  and 
ferric  chloride.  The  gold  so  obtained  is  extremely  finely  divided,  and 
an  addition  of  lead  acetate  and  sulphuric  acid  is  necessary  to  obtain  a 
sufficiently  dense  precipitate.  The  gold  precipitate  is  smelted  with  lead 
in  much  the  same  as  the  copper  precipitate.  The  wasted  liquor  from 
the  copper  extraction  is  worked  up  for  ferrous  sulphate,  which  by  roast- 
ing, gives  a  red  ferric  oxide,  which  is  used  for  paint. 

The  annual  production  of  the  works  is  as  follows : 

Copper  sulphate,  1,600  tons 

Ferrous  sulphate,  300  tons 

Iron  oxide,  red  paint,  1,000  tons 

Silver,  400  kilograms 

Gold,  100  kilograms 

The  Bradly  Process.' — In  the  Bradly  process  the  sulphide  ore  is 
carefully  roasted,  in  what  he  calls  an  amphidizer,  which  consists  of  a 
rotary  drum  with  a  central  heating  flue  through  which  heat  may 
be  supplied.  The  rotation  of  the  drum  operates  to  conduct  the  ore 
and  air  through  the  drum  in  one  direction  and  return  it  in  another 
to  the  same  end,  so  that  the  copper  is  more  or  less  completely  sul- 
phated   and   the  iron   oxidized.     The  roasting  is  conducted  at  a  tem- 

^E.  andM.  J.,  Jan.  6,  1912:  U.  S.  Patent  1,011,562,  Dec.  12,  1911. 


perature  of  from  450  to  550°  C,  air  being  blown  into  the  apparatus  to 
hasten  the  oxidation. 

The  roasted  ore  is  then  treated  with  an  excess  of  calci-um  chloride 
solution  in  a  reaction  drum,  while  the  temperature  is  maintained  at 
about  100°  C.  Cupric  chloride  is  produced  by  the  reaction  between  the 
copper  sulphate  and  calcium  chloride,  while  any  ferric  sulphate  in  the 
ore,  due  to  the  roasting,  reacts  with  calcium  chloride  to  produce  ferric 
chloride.  The  calcium  sulphate  resulting  from  both  these  reactions  is 
insoluble  and  is  separated  by  filtration  in  the  succeeding  step.  The 
production  of  ferric  chloride  at  this  point  is  advantageous  in  that  it  dis- 
solves copper  oxide,  copper  sulphide  or  metallic  copper,  which  remained 
unaffected  by  the  roasting,  producing  copper  chloride,  and  this  ferric 
chloride  also  maintains  the  copper  chloride  in  the  cupric  condition. 

The  gold  and  silver  in  the  ore  are  brought  into  solution  by  convert- 
ing all  the  copper  into  cupric  chloride  and  then  adding  a  small  amount 
of  chlorine,  chlorous,  or  chloric  compounds.  The  chlorides  of  silver 
and  gold  being  soluble  in  calcium  chloride  solutions  may  afterward  be 
precipitated  with  the  copper  and  subsequently  separated.  After  leav- 
ing the  reaction  drum  the  mass  of  gangue,  solution,  and  precipitates  is 
subjected  to  filtration.  The  solid  matter  forms  a  cake  which  consists  of 
the  gangue  in  the  ore  except  a  small  amount  of  iron  and  alumina  which 
have  l^een  taken  into  solution  and  the  calcium  sulphate  precipitate  already 
mentioned.  The  solution  comprises  a  carrier  in  which  has  been  dissolved 
the  metals  to  be  recovered,  a  small  amount  of  iron  and  alumina  and  any 
zinc  which  may  have  been  in  the  ore;  the  arsenic  will  have  been  separated 
by  filtration,  as  it  has  been  rendered  insoluble.  The  solution  is  then 
subjected  if  necessary  to  a  further  oxidizing  operation  in  order  to  be 
sure  that  the  metals  are  all  combined  at  their  highest  valency. 

The  solution  is  then  in  condition  for  treatment  for  the  separation 
of  the  dissolved  metals.  The  precipitation  of  iron  and  alumina  may  be 
made  by  cupric  oxide,  hydrate  or  calcium  carbonate,  and  as  this  precipi- 
tate will  carry  some  copper  it  is  returned  to  the  amphidizer,  or  roasting 
furnace,  after  having  been  removed  from  the  solution  by  filtration.  In  the 
amphidizer  the  iron  and  alumina  in  the  precipitate  are  rendered  insoluble, 
while  the  copper  is  left  in  a  soluble  condition  and  can  be  recovered.  The 
solution  from  which  the  iroii  and  alumina  has  been  removed  and  which 
then  contains  the  bulk  of  the  copper  is  run  into  a  second  tank  in  which  the 
copper  is  precipitated  by  carbonate  of  lime  as  oxide  of  copper.  This 
precipitate  is  filtered  from  the  solution  and  the  copper  is  recovered  by 
further  treatment  such  as  by  reduction  in  an  ordinary  smelting  furnace. 

Any  silver  and  gold  in  the  solution  is  carried  down  during  the  pre- 
cipitation of  the  iron,  aluminum,  and  copper,  and  finally  recovered  by 
separation  from  the  latter  metal.  Zinc  contained  in  the  ore  passes  into 
solution  as  chloride  of  zinc  and  accumulates.     It  is  therefore  necessary 




Dust  and  Fumes 


Heater  and   (— 
Asritatinff  ~^ 


Reaction  Drum 


Heating:  Coil 

Fig.  48. — Bradly  process.     Diagrammatic  sketch. 


at  stated  times  to  run  the  solution,  or  a  part  of  it,  after  the  final  treatment 
and  before  returning  it  to  the  reaction  drum,  to  a  third  precipitator  in 
which  the  zinc  is  precipitated  by  means  of  caustic  lime.  The  regenerated 
solution  from  which  the  gangue  and  all  metallic  compounds  has  been 
removed  and  which  contains  calcium  chloride  is  returned  to  the  reduction 
drum  for  the  treatment  of  additional  ore  from  the  amphidizer,  thus 
completing  the  cycle. 

During  the  roasting  considerable  dust  and  fumes  are  given  off,  con- 
taining sulphurous  and  sulphuric  anhydride.  This  is  condensed  by  a 
portion  of  the  solution  diverted  from  the  main  stream,  after  filtration 
from  the  ore,  and  is  again  returned  to  the  reaction  drum. 

Fig.  48  shows  a  diagrammatic  sketch  of  the  Bradly  process. 

Longmaid -Henderson  Process  for  Treating  Pjrritic  Cinders. — (The 
descriptions  of  the  early  work  of  this  process,  and  of  European  practice, 
are  principally  by  Clapham,  Wedding,  Ulrich,  Gill  and  Lunge;  the 
descriptions  here  given  is  taken  largely  from  Lunge's  Treatise  on  the 
Manufacture  of  Sulphuric  Acid  and  Alkali,  Vol.  1, 1891.) 

Sulphuric  acid  is  largely  made  from  pyritic  ores,  many  of  which  contain 
copper  worth  recovering,  after  the  sulphur  has  been  roasted  off  as  sulphur 
dioxide  for  the  sulphuric  acid  works.  Such  ores  are  quite  widely  scat- 
tered, but  the  largest  and  most  valuable  deposits  occur  in  Spain.  From 
the  Rio  Tinto  mines  in  Spain  the  ore  is  shipped  to  many  of  the  commer- 
cial and  manufacturing  nations  of  the  world.  During  the  year  1874, 
there  existed  in  Great  Britain  alone  22  copper  works  in  which  450,000 
tons  of  pyritic  cinders  were  treated  annually  by  wet  methods  for  the  ex- 
traction of  the  copper.  Two  of  these  works  made  sulphate,  three  pro- 
duced refined  copper,  and  the  rest  sold  their  cement  copper  precipitate 
to  copper  refineries.  In  1882  the  quantity  of  pyritic  cinders  treated  for 
copper  in  Great  Britain  amounted  to  434,427  tons,  containing  15,300 
tons  of  copper. 

The  process  used  in  England  at  that  time  is  still  largely  in  use  to-day, 
both  in  Europe  and  the  United  States,  and  with  very  little  modification 
from  the  original  process. 

The  Longmaid-Henderson  Process  consists  essentially  of  roasting  the 
pyritic  cinders,  from  the  sulphuric  acid  works,  with  salt,  leaching  out 
the  chloride  of  copper  so  formed,  followed  by  precipitation  with  iron. 
Longmaid  obtained  a  patent  dated  October  20,  1842,  and  another  in 
January,  1844,  both  relating  to  the  treatment  of  pyritic  cinders  by  roast- 
ing with  salt.  Longmaid  described  the  principles  of  the  process  very 
much  as  it  is  carried  out  to-day,  certainly  with  a  view  of  making  salt  cake 
and  chlorine  as  the  principal  products,  and  he  worked  it  out  on  a  large 
scale;  so  that  he  must  be  regarded  as  the  founder  of  the  wet  extraction 
of  copper.  Gossage,  in  1850,  first  employed  sponge  iron  for  precipitating 
the  copper.     Henderson  carried  the  process  to  greater  perfection.     In 



1865  he  erected  a  plant  at  Hebburn  for  the  Bede  Metal  Company,  to 
extract  copper  from  pyritic  cinders,  the  process  of  which  he  had  pro- 
tected by  patent.  Henderson's  principal  improvement  was  the  intro- 
duction of  absorption  towers  through  which  the  acid  gases  from  the 
chloridizing  roasting  were  condensed  and  yielded  a  weak  acid  solution, 
which  was  then  used  in  leaching  the  copper. 

The  most  important  ores  treated  by  this  process,  in  English  works, 
were  from  Spain  and  Portugal,  and  contained  from  47  to  49  per  cent. 
sulphur;  from  3.50  to  3.80  per  cent,  copper  and  from  0.75  to  1.20  oz.  in 
silver,  per  ton. 

In  the  German  works,  at  Oker,  the  ordinary  copper  ores  treated  by 
this  method  contained  60  per  cent,  iron  pyrites,  23  per  cent,  copper 
pyrites,  6  per  cent,  blende,  2  per  cent,  galena,  and  9  per  cent,  gangue. 

The  steps  in  the  process  may  be  summarized  as  follows: 

1.  Mixing  the  ore  with  salt  and  then  grinding  the  mixture. 

2.  Chloridizing  roasting. 

3.  Leaching  the  roasted  ore. 

4.  Precipitating  the  silver  from  the  argentiferous  liquors. 

5.  Precipitating  the  copper  from  the  desilverized  liquors. 

6.  To  which  may  be  added,  the  preparation  of  the  residue,  rich  in 
iron,  for  the  iron  smelters.     This  is  usually  done  at  the  copper  works. 

The  percentage  of  sulphur  in  cinders  as  supplied  by  the  acid  works 
to  the  copper  extraction  plants  varies  from  2  to  10  per  cent.  A  fair 
average  may  be  considered  from  4  to  5  per  cent.  At  Oker  the  cinders 
contained  from  5  to  8  per  cent,  sulphur  and  from  6  to  9  per  cent,  copper. 

The  following  analyses  by  Gibb  shows  the  composition  of  the  pyritic 
cinders  as  the  copper  works  received  them: 

Rio  Tinto 


San  Domingo 



Iron             ■  Calculated  as  CU2S  and  Fe2S3. . . 





0  39 

Zinc  oxide 

6  46 

0  06 


Cobaltic  oxide 

Calcium  oxide 

2  30 


Sulphuric  acid 

6  56 









Philips  gives  the  following  as  the  composition  of  cinders  from  San 
Domingo  ore: 











Oxygen  and  loss. 


58.25  (83.0  FC2O3) 








Samples  from  Widnes  and  Hebburn  showed, 

Widnes  Hebburn 

Copper,                                                  4 .  08  per  cent.  5 .  75'  per  cent. 

Sulphur,                                                    4. 12  per  cent.  3. 75 per  cent. 

Of  this  there  was  soluble  in  water, 


Soluble  in  hydrochloric  acid, 



Copper  insoluble  in  water  and  HCl, 

The  ores  that  used  to  be  treated  at  Oker,  showed,  on  an  average  of  a 
month's  run: 

46.0  per  cent.     26.1  per  cent. 
43.0 per  cent.     37.0 per  cent. 

22. 2  per  cent.  13. 3  per  cent. 
55 . 0  per  cent.  59 . 0  per  cent. 
31. 8  per  cent.     60. 6  per  cent. 

Copper  (principally  as  CuO), 

Iron  (principally  as  FcjOj), 

Lead  (as  PbO), 


Zinc  (as  ZnO), 

Manganese  (as  MujOJ, 


Sulphuric  acid  (corresponding  to  3.8  per  cent. 


Other  gangue. 


per  cent, 
per  cent. 

2.09  per  cent. 
0.008  per  cent. 
1 .  95    per  cent. 

per  cent. 

per  cent. 

per  cent. 

per  cent. 

per  cent. 


The  pyritic  cinders  as  received  by  the  copper  extraction  works, 
were  first  finely  ground  to  about  8  or  10  mesh  and  at  the  same  time 
mixed  with  a  sufficient  quantity  of  salt.  In  hand  furnaces  the  amount 
of  salt  varied  from  10  to  20  per  cent,  but  in  mechanical  furnaces  it  was 
less — about  7  1/2  per  cent.  At  Oker  the  cinders  were  mixed  with  15  per 
cent,  carnallite,  which  contains  chlorides  of  magnesium,  potassium,  so- 
dium, and  calcium;  all  of  which  assist  in  the  chloridizing  roasting. 

The  chloridizing   was   done   in  reverberatory,   and   in   muffle  fur- 


naces.  Sometime  in  a  sort  of  combination  between  the  two,  in  which 
only  a  part  of  the  furnace  was  muffled.  See  Figs.  10,  11,  and  12,  pages 
98  and  99. 

The  chloridizing  roasting  of  these  cinders  is  in  detail  essentially 
the  same  as  described  in  the  chapter  on  chloridizing  roasting.  Whatever 
the  means  of  furnace  employed,  the  object  to  be  attained  is  to  sufficiently 
convert  the  copper  into  sulphate,  which  owing  to  the  presence  of  the 
chloride  salts,  at  once  forms  with  the  cupric  sulphate,  by  mutual  decom- 
position, cupric  chloride  and  sodium  sulphate.  The  iron  should  be  con- 
verted as  completely  as  possible  into  the  ferric  oxide,  so  as  to  be  insol- 
uble both  in  water  and  dilute  acids. 

In  roasting,  the  SOg  and  0  acting  upon  NaCl,  chlorine  is  evolved 
which  greatly  aids  in  chloridizing  the  copper  as  well  as  any  other  metals 
present.  At  the  same  time  a  large  quantity  of  hydrochloric  acid  is 
formed  which  converts  the  oxides  of  copper,  silver,  zinc,  etc.,  into  chlo- 
rides, while  at  the  temperature  of  the  roasting  furnace  ferric  chloride 
cannot  exist.  Since  chlorides  of  copper  are  both  unstable  and  volatile 
at  very  high  temperatures,  a  low  red  heat  ought  not  to  be  exceeded; 
so  that  any  copper  pyrites  still  present  in  the  cinders  is  not  burned,  and 
therefore  escapes  chloridization. 

Wedding  describes  in  detail  the  roasting  at  Widnes  as  carried  on  in 
a  gas  furnace.  The  charge  of  4500  lb.  of  ore,  mixed  with  17  per 
cent,  salt,  is  spread  out  evenly  on  the  hearth  and  slowly  heated  till  a 
low  red  heat  has  been  reached  nearest  the  fire  bridge;  the  charge  is 
rabbled  and  left  to  itself,  the  gas  being  shut  off  but  the  air  allowed  to 
enter,  so  that  after  2  hours  scarcely  any  glowing  can  be  perceived 
at  the  fire  bridge.  After  1  hour's  and  3  hours'  roasting,  respect- 
ively, the  copper  of  the  charge  behaved  as  follows: 

1  hour's  roasting  3  hours'  roasting 
Soluble  in  water,                             54  per  cent.  51  per  cent. 

Soluble  in  hydrochloric  acid,         38  per  cent.  42  per  cent. 

Soluble  in  nitric  acid,  8  per  cent.  7  per  cent. 

After  3  hours  the  charge  is  quite  dark,  and  is  now  well  rabbled. 
There  ought  to  be  no  necessity  for  more  fire,  as  the  temperature  from 
first  should  have  been  raised  to  the  proper  point.  The  charge  is  now 
rabbled  regularly  at  short  intervals,  and  the  temperature  of  itself  rises 
in  consequence  of  the  chemical  reactions.  The  rise  becomes  sensible 
after  4  3/4  hours,  counting  from  the  beginning;  so  that  after  5  1/4  hours 
a  dark  red  heat  is  reached.  Up  to  this  point  there  is  a  copious  evolution 
of  white  vapors  and  blue  flames;  from  this  period  there  is  less  of  these, 
and  it  is  the  roasterman's  principal  task  to  see  that  heating  of  the  charge 
is  uniform,  and  that  some  places  do  not  show  more  flame  than  others. 
After  6  1/2  hours  these  flames  are  almost  entirely  gone;  and  this  fact, 


along  with  the  greenish-gray  color  of  the  charge,  are  the  principal 
tests  for  j  udging  whether  the  operation  is  finished.  A  sample  is  now  taken, 
and  if  its  examination  shows  the  completion  of  the  roasting  process, 
the  charge,  which  has  now  been  6  1/2  or  6  3/4  hours  in  the  furnace,  is 
withdrawn.     Of  the  copper  now, 

75  per  cent,  is  soluble  in  water, 
20  per  cent,  is  soluble  in  hydrochloric  acid, 
5  per  cent  is  soluble  in  nitric  acid. 

In  roasting,  in  order  to  get  the  best  results,  the  charge  is  first  heated 
to  a  temperature  sufficiently  high  to  start  the  chemical  reactions,  and 
then  to  maintain  these  at  the  lowest  possible  temperature  up  to  the 
finish,  and  to  see  that  the  entire  mass  of  ore  is  uniformly  treated. 

It  is  of  great  importance  not  to  leave  the  ore  any  longer  in  the  fur- 
nace than  necessary  after  the  roasting  is  completed.  The  depth  of  the 
ore  is  usually  from  4  to  5  in.,  and  this  depth  of  charge,  while  more  diffi- 
cult to  rabble,  facilitates  the  chloridization,  since  the  gas  rising  in  the  ore 
heated  both  at  top  and  bottom  has  all  the  more  opportunity  of  coming 
in  contact  with  all  parts  of  it. 

At  Oker,  where  gas-fired  furnaces  exactly  like  those  at  Widnes  were 
used,  each  charge  of  5000  lb.  of  ore  with  15  per  cent,  carnallite  is  brought 
to  a  low  red  heat  in  4  hours;  the  firing  is  then  discontinued  and  the 
mass  rabbled,  for  4  or  5  hours  with  a  very  low  fire  or  without  any,  the 
air  valves  being  meanwhile  kept  open  in  order  to  allow  the  air  to  act  on 
the  charge.  The  charge  is  then  withdrawn  and  a  fresh  charge  introduced 
as  soon  as  the  furnace  is  empty.  Five  tons  of  roasted  ore  were  worked 
off  in  24  hours,  with  a  coal  consumption  of  10  to  12  per  cent.  The  com- 
position of  the  roasted  ore  after  chloridizing  roasting  with  20  per  cent, 
carnallite  was  as  follows: 

Soluble  in  Wateb 










K^SO^  20.50 



■  cent. 

Per  cent. 

3 .  86    calculated  as  CuCl^, 


0 .  005  calculated  as  AgCl, 


0 .  60    calculated  as  FeClj, 


0.17    calculated  as  Al^CSO J ,, 


1 .  64    calculated  as  ZnClj, 


0 .  75    calculated  as  MnClj, 


0 .  07    calculated  as  NiCl^, 


1 .  60    calculated  as  CaClj, 














Insoluble  in  acids, 

Insoluble  in  Water 
Per  cent. 

2.57    calculated  as 
1.17    calculated  as 

34 .  56    calculated  as 








calculated  as 
calculated  as 

calculated  as     CaSO,, 

Per  cent. 




In  mechanical  furnaces  less  salt  was  used  than  in  hand  roasters — 
averaging  7  1/2  per  cent,  as  against  15  per  cent,  in  the  hand  furnaces. 
Sometimes  only  a  portion  of  the  salt  was  added  at  the  start,  and  the 
remainder  added  afterward.  In  the  mufHe  roasters  the  ore  is  first 
roasted  9  hours  with  12  per  cent,  salt  and  another  3  hours  with  8  per 
cent,  more  salt.  In  the  combination  furnaces  with  protecting  arch  the 
weight  of  the  ore  was  5800  lb.,  and  the  time  of  roasting  8  hours;  in  the 
mechanical  furnaces  5  tons  were  roasted  in  9  hours. 

Gibb  investigated  the  comparative  working  of  the  different  furnaces, 
which  may  be  summarized  as  follows: 

Cupric  chloride.  . 
Cuprous  chloride. 
Cupric  oxide. 
Sodium  chloride. . 
Sodium  sulphate . 
Insoluble  copper. 

Gas  furnace 

Muffle   furnace 

Per  cent. 

4.03  = 

Cu.  per 

Per  cent. 

Mechanical  fur. 




4,25  = 










Cu,  per   _ 

^       Per  cent, 

6.70  = 





Cu,  per 






The  principal  object  in  roasting,  of  course,  is  to  get  as  much  copper 
as  possible  soluble  in  water  or  dilute  acids.  In  the  above  comparison 
there  is  a  slightly  better  result  in  favor  of  the  muffle  roaster,  but  not 


enough  to  give  a  decided  advantage,  and  the  advantage  is  largely  over- 
balanced by  the  increased  fuel  consumption  of  the  mufHe  furnaces. 

At  Oker  the  average  results  of  the  constantly  taken  samples  of  the 
chloridized  ore  showed  75  per  cent,  of  the  copper  was  soluble  in  water 
as  the  cupric  chloride  and  neutral  sulphate;  20  per  cent,  was  soluble  in 
dilute  hydrochloric  acid,  as  cuprous  chloride  and  oxychloride,  and  5  per 
cent,  was  insoluble  in  the  treatment  of  the  ore  for  the  copper  but  was 
soluble  in  aqua  regia. 

In  roasting,  the  sulphur  in  the  pyritic  cinders  must  bear  a  certain 
proportion  to  the  copper.  With  a  4  per  cent,  copper  ore  the  sulphur 
should  not  exceed  6  per  cent. ;  an  equal  percentage  of  sulphur  and  copper 
is  preferable.     If  less  sulphur  is  present,  raw  pyrites  must  be  added. 

The  test  for  ascertaining  the  cpmpletion  of  the  roasting  are  made  by 
taking  a  certain  definite  quantity  of  the  roasted  ore,  leaching  it  with 
water  and  dilute  hydrochloric  acid  just  as  in  the  regular  leaching;  the 
residue  is  then  boiled  with  aqua  regia,  supersaturated  with  ammonia, 
and  allowed  to  settle;  the  more  or  less  blue  color  of  the  ammonia-cupric 
salt  gives  a  sufficient  indication  of  the  percentage  of  insoluble  copper. 

Condensation  of  the  Furnace  Gases. — -In  all  but  the  mufHe  furnaces 
the  gas  from  the  roasting  ore  is  mixed  with  gas  from  the  fire-boxes. 
Even  in  the  muffie  furnaces  the  gases  from  the  ore  are  mixed  with  air 
to  such  an  extent  that  a  condensation  of  strong  acid  is  not  possible. 
The  acid  from  the  mufHe  furnaces  is  only  slightly  more  concentrated 
than  from  the  reverberatories;  but  this  is  not  a  serious  matter  as  the 
acids  are  always  used  in  a  very  dilute  solution  for  leaching.  The  furnace 
gases  contain  principally,  besides  oxygen  and  nitrogen,  sulphur  dioxide, 
sulphur  trioxide,  hydrochloric  acid,  chlorine,  and  very  small  quantities 
of  metallic  chlorides.  Henderson  proposed  volatilizing  the  copper 
entirely  as  cupric  chloride  and  condense  the  latter  in  towers;  but  this  has 
turned  out  quite  impracticable.  The  quantity  of  copper  volatilized  in 
the  ordinary  process  of  roasting  is  not  large,  about  1/4  per  cent,  of  the 
whole,  and  this  is  condensed  with  the  acid  in  the  tower  acid  for  leaching 
the  ore. 

The  condensation  of  the  gases  from  the  roasters  takes  place  in 
towers  made  of  brickwork  set  in  tar  and  sand  (or,  better,  of  stone  flags), 
and  packed  with  coke,  fire  bricks,  and  the  like.  Large  stoneware  pipes 
are  sometimes  employed.  Coke  can  be  used  for  the  filling  material  with 
the  muffle  roasters;  but  the  other  furnaces  require  brick,  or  similar 
material  and  must  have  larger  condensers,  as  these  towers  have  to  serve 
for  a  larger  volume  of  gas.  The  size  of  the  towers  varies  with  the  plant; 
for  12  furnaces  a  tower  of  8  ft.  square  and  40  to  50  ft.  high  is  sufficient. 
The  gas  enters  at  the  bottom,  meets  a  spray  of  water  coming  from  the 
top,  which  washes  the  acid  out  of  it,  and  again  leaves  the  tower  at  the 
top,  whence  it  is  taken  downward  into  a  flue  leading  to  the  chimney. 


The  total  condensed  liquid,  which  is  a  mixture  of  sulphuric  and 
hydrochloric  acids,  is  used  in  the  succeeding  operation  of  leaching,  and 
frequently  is  not  even  sufficient  for  dissolving  all  the  copper  oxide  and 
cuprous  chloride.  The  sulphur  dioxide  in  the  liquid  is  oxidized  to  sul- 
phuric acid  by  the  action  of  the  chlorine. 

Leaching  of  the  Roasted  Ore. — The  roasted  ore  is  carried  in  bogies  on 
tramways  to  the  leaching  tanks.  The  material  used  in  the  construction 
of  these  tanks  is  wood.  It  is  most  difficult  to  prevent  leakage  in  large 
wooden  tanks  when  using  a  hot  chloride  solution;  on  this  account  the 
entire  floor  of  the  leaching  shed  is  covered  with  a  thick  layer  of  asphalt 
and  slopes  to  one  side,  so  that  all  liquors  Can  be  recovered  in  a  catch-well. 

The  leaching  tanks  are  square,  about  11X11  ft.  wide,  and  4  to  5  ft. 
deep,  made  of  well  seasoned  and  planed  3-in.  planks,  secured  by  corner 
pieces,  screw-bolts,  etc.  The  joints  are  tightened  by  putting  a  little 
redlead  between  the  planks  before  putting  them  together;  the  bottom 
joints  are  best  caulked  with  tarred  spun  yarn,  and  the  whole  tank  painted 
with  hot  coal  tar.  At  Oker  lead-lined  tanks  were  used,  but  they  were 
very  expensive  and  needed  frequent  repairing.  On  the  bottom  of  the 
tank  are  placed  slats,  laid  on  end;  and  upon  these  perforated  tiles  or 
boards;  upon  this  false  bottom  a  layer  of  sifted  furnace  cinders  is  spread 
out,  and  on  top  of  this  a  layer  of  sand,  coke,  or  straw  from  3  to  6  in.  deep. 

The  ore  is  charged  upon  the  bed  so  prepared  and  is  then  ready  for 
leaching.  The  leach  liquors  are  conveyed  in  earthenware  and  india- 
rubber  tubes  of  3  to  4  in.  in  diameter,  which  are  provided  with  iron 
pinch-clamps.  In  order  to  force  the  liquors  from  one  tank  to  the  other, 
or  from  the  catch- well  into  the  tanks,  simple  stoneware  injectors  are 
provided.     Each  tank  has  a  steam  pipe  for  heating. 

In  each  tank  there  is  put  10  tons  of  roasted  ore,  quite  hot,  from 
the  furnace,  and  is  covered  with  a  weak  liquor  from  a  previous  operation, ' 
which  gets  heated  by  the  heat  of  the  mass  itself.  After  one  or  two 
hours  the  now  concentrated  liquor  is  run  off  by  a  plug-hole  below  the 
false  bottom,  and  is  delivered  to  the  precipitating  tanks.  The  plug  is 
put  in  again,  and  the  ore  leached  with  hot  water;  thus  weaker  liquors 
are  produced  which  are  forced  to  another  tank,  as  described.  Generally 
three  waters  are  put  on,  and  thus  most  of  the  purest  copper  and  95  per 
cent,  of  all  the  silver  contained  in  the  pyrites  are  obtained.  After  the 
treatment  with  the  water  the  dilute  acid  solution  from  the  tower  is 
applied,  sometimes  as  many  as  six  applications  before  all  the  copper  is 
satisfact orily  extracted. 

The  liquors  obtained  by  use  of  acid  solutions  contain  many  impuri- 
ties, especially  arsenic,  bismuth,  antimony,  and  lead — according  to  Gibb, 
for  each  100  parts  of  copper,  5.4  arsenic  and  0.3  bismuth;  these  liquors 
are  usually  treated  separately  in  most  works  because  they  yield  impure 



As  a  rule  each  solution  application  is  allowed  to  stand  only  a  few 
hours  on  the  ore;  the  nine  washings  of  each  tank,  together  with  the 
filling  and  emptying,  takes  about  48  hours. 

The  effect  of  the  leaching  is  best  seen  from  the  following  by  Gibb, 
which  at  the  same  time  shows  the  difference  in  work  between  mechanical 
and  hand-worked  furnaces: 

Soluble  in  water 

Cupric  chloride, 
Cuprous  chloride, 
Cupric  sulphate. 
Ferrous  sulphate, 
Ferric  sulphate, 
Zinc  sulphate, 
Calcium  sulphate, 
Sodium  sulphate. 
Sodium  chloride. 

Mechanical  furnace 
per  cent. 
4.16  =  1.96  per  cent.  Cu 

1.83  0.80 

19.36  2.77 

Soluble  in  dilute  HCl 

Cuprous  chloride,  0.15   =0.01  per  cent.  Cu 

Cupric  oxide,  0.225    0.18 

Residue  by  difference,  80 . 40      0. 08 

100.00      3.04 

Sodium  chloride  equiva- 
lent to  sodium  salts 
as  above,  7.56 

Hand-worked  furnace 
per  cent. 
3 .  81  =  1 .  82  per  cent.  Cu 
0.19  0.12 

21.11  1.94 

0.33  =  0.21  per  cent.  Cu 
1.01     0.81 
77.55     0.11 

100.00     3.07 


At  Oker,  the  process  is  carried  out  as  follows:  The  roasted  ore  is 
leached  in  charges  of  5  tons  each,  first  with  the  final  liquor  of  a  previous 
charge,  100  parts  of  this  liquor,  of  1.145  specific  gravity,  contained: 















CoO;  NiO, 












As;  Sb, 


Total  solids,  14.49  per  cent. 

This  liquor,  already  heated  in  pumping  by  the  injector  to  50°  C.  is 
further  heated,  when  it  comes  in  contact  with  the  hot  roasted  ore  to 
nearly  the  boiling-point.  When  the  charge  is  thoroughly  saturated 
with  the  liquor,  the  spigot  is  opened  and  the  liquor  allowed  to  drain  as 
long  as  it  shows  any  blue  color.     This  lasts  from  4  to  5  hours  and  furn- 


ishes  a  copper  liquor  of  1.355  specific  gravity  and  of  the  following  com- 


















1  0.29 











Total  solids, 






After  the  first  leaching  is  over,  the  dilute  condenser  acid,  first  brought 
to  boiling,  is  run  into  tanks  and  allowed  to  act  for  24  hours;  then  it  is 
drawn  off,  and  a  third  leaching  effected  by  sulphuric  acid.  For  5  tons 
of  ore  250  lb.  of  chamber-acid  of  106°  Tw.,  diluted  to  12°  Tw.  and  heated 
to  boiling,  is  employed  and  allowed  to  remain  in  contact  with  the  ore 
for  two  days,  or  until  the  liquor  acquires  a  neutral  reaction. 

The  first  copper  liquors  contain  most  of  the  silver,  and  are  therefore 
kept  apart  from  the  later  liquors,  which  do  not  contain  as  much  silver. 

The  cupric  chloride  is,  of  course,  easily  dissolved  in  the  final  liquor; 
the  cuprous  chloride  in  the  presence  of  alkaline  chlorides  is  also  dissolved 
at  a  higher  temperature  without  difficulty;  lastly  cupric  oxide  is  to  be 
converted  into  cupric  and  cuprous  chlorides  by  the  ferrous  chloride  of 
the  final  liquors  from  previous  applications: 

2FeCl2  +  3CuO  =  Feft^  +  CuCl^  +  Cu^Cl^ 

but  this  could  only  be  done  by  an  intimate  mechanical  mixture  of  the 
liquor  with  the  ore;  and  it  is  therefore  preferred  to  dissolve  merely  75  to 
80  per  cent,  of  the  copper  by  means  of  the  final  liquor,  and  the  remainder 
by  further  leaching  with  dilute  acids. 

Precipitation  of  the  Copper. — The  precipitation  of  the  copper  is  some- 
times preceded  by  a  special  treatment  for  obtaining  the  silver,  separately 
from  the  copper. 

The  precipitation  of  the  copper  is  universally  done  by  means  of 
metallic  iron.  Gibb  proposed  precipitating  the  copper  with  hydrogen 
sulphide,  but  this  method  was  later  given  up  at  the  Bede  Metal  Works 
where  it  was  tried  on  a  large  scale. 

For  precipitation  with  iron,  the  iron  used  was  either,  wrought  scrap, 
or  sponge  iron  reduced  from  the  residues.  Light  scrap  is  better  than 
heavy  scrap,  but  the  copper  precipitated,  owing  to  more  impurities, 
gives  a  lower  grade  cement  copper.  The  precipitation  takes  place  in 
wooden  tanks  like  those  used  in  leaching  the  ore,  and  are  furnished  with 
a  steam  pipe  for  heating  the  copper  liquor.  The  tanks  are  filled  with 
scrap  iron;  copper  liquor  is  run  upon  it  and  the  steam  turned  on.  The 
heating  is  continued  till  a  bright  strip  of  iron,  held  in  the  liquid,  no 



longer  indicates  the  presence  of  copper  in  solution.  At  Oker,  according 
to  the  degree  of  concentration  of  the  liquors,  the  boiling  takes  place 
two  or  three  times  before  all  the  copper  is  thrown  down;  the  process 
lasts  from  one  to  three  days,  and  requires  as  much  iron  as  the  weight  ot 
copper  produced,  which  proves  that  a  large  part  of  the  copper  must 
have  been  in  solution  as  cuprous  chloride.  Once  a  month  the  precipi- 
tated copper  is  removed  from  the  tanks  and  washed.  If  sponge  iron 
is  used  for  precipitation  of  the  copper,  continuous  stirring  is  required, 
for  which  at.  some  works  mechanical  agitators  are  used,  at  others  manual 
labor.  At  the  Bede  Metal  Works  an  india-rubber  hose,  through  which 
a  blast  of  air  passes,  is  moved  about  in  the  tank.  Perfect  mixture  is 
thus  obtained,  and.  the  precipitated  copper  contains  only  1  per  cent, 
metallic  iron. 

The  composition  of  the  copper  precipitated  by  the  various  methods 
is  shown  by  the  following  analyses  by  Gibb : 

Precipitated  by 





Ferric  oxide 



Sponge  iron 

Per  cent. 

Heavy  scrap 

Per  cent. 

Light  scrap 

Per  cent. 

At  Oker  the  composition  of  the  copper  precipitated  by  scrap  iron 
and  dried  at  100°  G.,  was: 


CI  ' 

Insoluble  in  acids, 

Oxygen-moisture  (by  loss), 





































The  copper  precipitate  from  pyritic  cinders  made  at  the  Witkowitz 
works,  dried  at  100°  C,  was  composed  as  follows: 















30  1 


31  )  = 


































=  69.45  per  cent.  Cu 


The  cement  copper  from  the  leaching  works  may  be  smelted  to 
blister  copper,  or  sent  to  the  smelting  works.  The  copper  precipitated 
from  aqueous  solutions,  if  kept  separate  from  that  from  the  acid  solu- 
tions, can  be  smelted  direct  to  blister  copper  by  adding  to  it  lime  and 
slags;  the  copper  from  the  acid  solutions  is  frequently  so  impure  that  it 
has  to  be  mixed  with  raw  ore,  and  smelted  for  "coarse  metal"  which 
yields  blister  copper  only  after  second  treatment.  At  some  works  both 
precipitates  are  melted  together,  being  charged  into  the  furnace  while 
moist.  The  slag  produced  from  this  operation,  containing  from  3  to  10 
per  cent,  copper,  are  charged  into  blast  furnaces. 

The  furnace  for  smelting  the  copper  precipitate  used  at  English  wet- 
extraction  works  are  reverberatory  furnaces  of  the  well-known  Swansea 
type.  After  smelting,  the  slag  is  skimmed  off,  and  the  copper  tapped  as 
blister  copper.  When  sponge  iron  has  been  used,  the  excess  of  carbon 
prevents  the  copper  from  being  melted  directly  into  blister  copper; 
therefore  about  one-half  of  the  precipitate  roasted  in  large  furnaces 
similar  to  those  used  in  roasting  the  ore.  Here  the  carbon  is  burned 
off  and  the  copper  partly  oxidized;  the  roasted  precipitate  is  mixed  with 
raw  precipitate  and  smelted  for  blister  copper.  The  blister  copper  is 
refined  by  roasting  to  oxidize  the  iron,  sulphur,  etc.,  followed  by  reducing 
with  charcoal  the  oxide  of  copper  produced  in  the  roasting,  and  poling 
according  to  the  method  of  smelting  usually  employed  by  the  copper 
smelter.     The  copper  produced  in  this  way  is  pure  and  tough. 



Precipitation  of  the  Silver. — Most  cupriferous  pyrites  contain  some 
silver,  and  small  quantities  of  gold.  The  cupriferous  pyrites  cinders 
from  Spain,  according  to  Phillips,  contain  on  an  average  of  0.0027  per 
cent,  silver,  and  0.0001  per  cent.  gold.  On  roasting  with  salt,  most  of 
the  silver  and  much  of  the  gold  is  converted  into  the  chlorides.  Owing 
to  the  solubility  of  silver  chloride  in  a  solution  of  other  chlorides,  and 
the  solubility  of  gold  chloride  in  water,  the  silver  and  gold  chlorides  are 
extracted  with  the  copper  by  the  leaching  solvent.  At  the  present 
time,  when  copper  is  refined  electrolytically  at  a  small  cost,  no  attempt 
is  usually  made  to  recover  the  gold  and  silver  separate  from  the  copper. 
In  earlier  years,  however,  this  separation  was  desired,  and  various 
schemes  were  proposed  for  their  separate  recovery. 

Claudit  Process. — Of  the  methods  proposed  for  the  separate  recovery 
of  the  silver  and  gold,  that  devised  by  Claudit  was  quite  generally 

This  process  consists  in  precipitating  with  a  soluble  iodide  the  silver 
from  the  liquors  in  the  state  of  Agl,  silver  iodide,  which  is  quite  insol- 
uble in  chloride  solutions.  Only  the  first  liquors,  rich  in  silver,  are  sub- 
jected to  the  Claudit  process  for  the  precipitation  of  the  silver.  The 
other  solutions  are  returned  to  the  ore,  or  they  are  too  dilute  in  sUver 
to  make  its  recovery  profitable.  The  liquors,  from  the  leaching  tanks, 
and  before  precipitating  the  copper,  are  run  into  settling  tanks,  where 
they  are  completely  settled  and  their  silver  contents  accurately  estimated 
by  adding  to  a  certain  volume  of  muriatic  acid  and  a  solution  of  lead  ace- 
tate, and  afterward  potassium  iodide.  The  precipitate  is  collected  on 
a  filter,  washed,  dried  and  fused  with  a  flux  of  soda,  borax  and  finely 
pulverized  carbon.  The  lead  regulus  is  cupelled,  and  from  the  weight 
of  the  silver  thus  obtained,  that  contained  in  the  liquors  in  the  settling 
tanks  is  computed.  To  the  liquor  in  the  settling  tanks  is  then  added  a 
solution  of  potassium,  sodium,  or  zinc  iodide  of  known  strength,  so  that 
the  quantity  is  just  sufficient  to  precipitate  all  the  silver;  the  iodide 
solution  is  diluted  to  such  an  extent  that  it  amounts  to  one-tenth  the 
volume  of  the  liquid. 

The  reactions  for  the  precipitation  of  the  sUver  with  sodium  and  zinc 
iodides  are: 

AgCH-Nal  =NaCl-|-AgI. 
2AgCl  -F  Znl^  =  ZnCIj  +  2AgI. 

The  precipitated  iodide  of  silver  is  allowed  to  settle  for  about  48 
hours,  and  tested  in  the  laboratory  to  see  if  the  precipitation  has  been 
complete.  The  liquors  are  then  run  into  the  copper  precipitating 
tanks,  where  they  are  treated  in  the  usual  way  for  the  precipitation  of 
the  copper. 


The  quantity  of  iodide  employed  for  the  precipitation  is  much 
larger  than  that  corresponding  to  the  silver  present,  since  a  portion 
of  the  lead  is  also  thrown  down  as  PbClj.  The  silver  is  probably 
precipitated  before  the  lead,  but  a  fractional  precipitation  is  not 
possible,  so  that  necessarily  a  corresponding  excess  of  the  precipitant  is 

The  precipitate,  consisting  principally  of  Agl,  Pblj  and  PbSO<(which 
is  deposited  on  cooling  the  liquor),  is  well  washed  with  water;  and  after 
a  sufficient  quantity  of  it  has  been  collected,  it  is  treated  with  metallic 
zinc  and  hydrochloric  acid  or  with  sodium  sulphide; 

2AgI  +  Zn  =  Znl2  +  2Ag.  or 
2AgI  +  Na^S  =  2NaH-  Ag^S. 

Thus  the  Agl  and  Pblj  are  decomposed  completely,  the  PbSO^  partly, 
and  liquor  containing  zinc  or  sodium  iodide  is  obtained,  which  is  employed 
again  and  the  cycle  continued  indefinitely. 

After  two  or  three  days  the  precipitated  liquor  is  drawn  off,  and 
run  into  copper  precipitating  tanks  where  the  copper  is  precipitated,  in 
the  usual  way.  The  liquors  may  still  contain  2  to  3  milligrm.  of 
silver  per  liter.  The  tanks  are  furnished  with  two  outlets,  one  at  the 
bottom  and  the  other  about  8  inches  higher.  The  desilverized  solution 
is  drawn  off  the  upper  hole  while  the  precipitate  remains  undisturbed 
on  the  bottom.  New  liquor  from  the  leaching  vats  is  then  let  in  and 
the  procedure  repeated  until  there  is  a  sufficient  silver  slime  in  the  bot- 
tom that  it  has  to  be  recovered.  This  is  usually  done  every  month  or 
two,  depending  on  the  percentage  of  silver  in  the  ore. 

The  precipitate  obtained,  treated  with  zinc  and  hydrochloric  acid, 
ready  for  melting,  contains  from  3  to  10  per  cent,  silver  and  usually 
some  gold.     An  anlaysis  shows  the  following  composition: 

Ag,  5.95 

Au,  0.06 

Pb,  62.28 

Cu,  0.60 

ZnO,  15.46 

Fe,0,  1.50 

CaO,  1 .  10 

SO3,  7.68 

Insoluble,  1  •  75 

Oxygen  and  loss,  3 .  62 


About  two-thirds  of  the  silver  and  gold  originally  contained  in  the 
roasted  ore  is  recovered,  with  care,  by  the  Claudit   method.     Under 


normal  conditions,  from  10  to  15  per  cent,  of  the  iodide  is  lost  in  pre- 
cipitating the  silver. 

Disposition  of  the  Residues. — The  residues,  from  the  leaching,  make 
a  valuable  ore  of  iron,  and  is  known  as  "purple  ore"  or  "blue  biHy-" 
In  this  way  the  additional  profit  may  be  realized.  The  following  is 
the  composition  of  two  average  samples: 

Ferric  oxide,  90.61  95.10 

Copper,  0.15  0.18 

Sulphur,  0.08  0.07 


Lead  sulphate,  1.46  1.29 

Calcium  sulphate,  0 .  37  0 .  29 

Sodium  sulphate,  0.27              

Sodium  chloride,  0. 28             

Insoluble,  6.30  2.13 

99 .62  99 . 55 

Metalic  iron  63.42  66.57 

This  shows  an  excellent  quality  of  iron  ore,  entirely  free  from  phos- 
phorus, and  containing  but  little  sulphur.  Its  slight  percentage  of 
copper  does  no  harm.  The  lead  contained  in  the  pyrites  remains  behind 
in  the  residue  in  the  shape  of  sulphate,  and  injures  its  quality  as  an  iron 
ore.  Schaflfner  proposed  drenching  the  residues  with  a  solution  of  cal- 
cium chloride,  heated  to  about  40°  C,  and  acidulated  with  hydrochloric 
acid.  By  mutual  decomposition,  gypsum  and  lead  chloride  are  at  once 
formed,  which  remains  dissolved  in  the  acid  liquor.  This  is  run  off  and 
brought  in  contact  with  metallic  iron,  which  precipitates  the  lead  in  the 
metallic  state.  After  washing,  the  purple  ore  is  quite  free  from  lead  sul 
phate.  At  the  same  time  the  CaClj  dissolves  the  last  traces  of  copper 
and  silver  present  as  CujClj  and  AgCl;  these  are  precipitated  along  with 
the  lead.  It  should  be  noted  that  hydrogen  sulphide  fails  to  indicate 
the  lead  in  a  solution  of  calcium  chloride  acidulated  with  hydrochloric 

The  only  drawback  to  the  use  of  pyritic  residues  in  smelting  for  iron 
is  their  fineness,  which  militates  against  their  desirability.  Unsuccessful 
attempts  for  a  long  time  were  made  to  agglomerate  this  ore  but  without 
success.  At  present,  however,  this  is  cheaply  done  by  sintering,  or  by 
partially  fusing  it.  It  can  cheaply  be  sintered  by  mixing  with  it  from 
7  to  10  per  cent,  coal  or  coke  dust,  the  mixture  moistened  to  the  proper 
consistency,  placing  it  on  a  permeable  hearth  such  as  broken  limestone,  and 
then  applying  suction  (down  draft)  so  that  the  intense  ignition  of  the 
coal  dust  fuses  the  ore  and  agglomerates  it  into  a  strong  porous  mass, 
and  puts  it  in  ideal  condition  for  smelting.' 

'John  E.  Greenawalt,  U.  S.  Patent,  839,064,  Dec.  18,  1906. 


JOxporimcnts  made  in  Denver  by  this  metliod  gave  excellent  results 
both  as  to  cost  of  sintering  and  product  obtained.  These  experiments 
have  since  been  duplicated  in  large  installations  in  working  plants. 

In  some  works  the  residues  are  briquetted,  and  the  briquetts,  after 
being  dried,  are  subjected  to  a  high  temperature,  so  that  by  partial  fusion, 
they  become  coherent.  In  still  other  works  the  residue  is  sintered  and 
nodulized  by  passing  the  purple  ore  through  a  rotary  cylinder  similar 
to  those  used  in  revolving  roasters,  in  which  the  ore  is  heated  to  a  high 
temperature  by  means  of  powdered  coal  blown  in  at  the  discharge  end. 
This  product  varies  in  size  from  that  of  rice  to  that  of  walnuts,  and  forms 
hard  balls  more  or  less  thoroughly  sintered.  By  the  Grondal  method 
the  cinder  is  formed  into  briquetts  of  uniform  size  by  an  automatic 
plunger  press.  The  briquetts  are  loaded  on  flat  cars,  which  are  then 
slowly  pushed  through  a  channel  furnace  150  to  200  ft.  long,  where  they 
are  gi-adually  heated  until  they  arrive  at  a  zone  of  the  furnace  where  the 
temperature  reaches  2400°  F.  The  finished  product  is  hard  and  strong, 
but  porous,  so  that  it  is  well  suited  to  blast  furnace  work  and  open  hearth 

The  cost  for  nodulizing  or  briquetting  will  vary  from  $0.75  to  $1.25 
per  ton.     Costs  for  sintering  have  not  yet  been  established. 

Longmaid -Henderson  Process  at  the  Helsingborg  Copper  Works, 
Sweden.— In  preparing  the  tanks  for  the  ore,  standard  bricks  are  placed 
on  edge  on  the  bottom.  On  these  special  perforated  brick,  12  in.  square, 
are  laid  as  a  false  bottom  for  a  filter  composed  of  straw  or  screened  lump 
cinder,  iron  ore  or  something  similar,  on  top  of  which  the  ore  is  charged 
and  leveled.  The  tanks  are  built  elliptical  in  shape,  one  diameter  of 
which  is  10  ft.  and  the  other  6  ft.  The  inside  depth  is  4  ft.  and  the 
thickness  of  stave  5  in. 

After  leveling  the  charge,  acid  liquors  of  5  to  8°  B.  strength,  ob- 
tained in  previous  leaching,  are  let  into  the  tanks.  The  ore  is  still 
warm  from  the  roasting.  The  first  liquor  issuing  from  the  vats  is  very 
strong — 40°  B.  It  contains  about  5  per  cent,  copper  and  nearly  all  the 
silver.  It  is  transferred  immediately  to  the  silver  precipitation  de- 

When  the  strength  of  the  liquor  as  it  is  drawn  from  the  vats  has 
gone  down  to  22°  B.,  it  is  taken  to  the  copper  precipitators.  This  con- 
tinues until  the  strength  of  the  outgoing  liquor  has  gone  down  to  10°  B. 
The  incoming  liquor  is  shut  off  and  in  its  place  acid  from  the  towers  is 
let  into  the  tank.  The  acid  is  allowed  to  remain  in  the  tanks  two  hours, 
at  the  end  of  which  time  the  liquor  is  exchanged  for  fresh  acid,  which  in 
its  turn  remains  two  hours.  The  liquor  so  obtained  is  used  in  the  first 
application  for  freshly  filled  tanks. 

After  this  leaching,  the  ore  is  washed  with  water,  first  with  wash  water 
from  a  previous  operation  and  finally  with  fresh  water.     To  test  whether 


all  the  copper  is  extracted,  a  well-polished  iron  plate  is  immersed  in  the 
off  coming  solution;  if  there  is  no  coloring  of  copper,  the  leaching  is 

The  wash  water  ought  to  have  a  temperature  of  about  50°  C.  The 
time  of  leaching  is  usually  about  40  hours  for  a  10-ton  charge. 

Of  the  copper  in  the  roasted  ore  about  80  per  cent,  is  soluble  in  warm 

16  per  cent,  soluble  in  weak  hydrochloric  acid, 
4  per  cent,  insoluble,  and  remains  in  the  purple  ore. 
The  purple  ore  is  briquetted  and  sintered  and  then  smelted  for  iron. 
The   crude  purple  oie  contains-  60.6  per  cent,  iron  and  0.17  percent, 
sulphur;  the  fused  briquetts  contain  60.6  per  cent,  iron  and  0.023  per  cent, 

Longmaid -Henderson  Process  as  Carried  out  at  the  Works  of  the 
Pennsylvania  Salt  Manufacturing  Co.,  Natrona,  Pa. — A  description  of  a 
large  modern  plant,  treating  approximately  200  tons  of  pyritic  cinders  a 
day  by  the  Longmaid-Henderson  process  is  given  by  Joel  G.  Clemer.' 
The  method  of  operation  is  as  follows: 

The  pyritic  cinders  are  ground  dry  in  Chillean  mills,  to  about  20  mesh, 
and  mixed  during  the  operation  with  10  per  cent,  salt,  testing  high  in 
NaCl.  The  mixture  is  then  lifted  by  bucket  elevators  to  overhead  hoppers 
from  which  it  is  drawn  into  tram  cars,  weighed,  and  dumped  into  the 
furnace  hoppers,  or  pipes  set  into  the  floor  over  the  furnaces.  Should 
the  unburnt  sulphur  in  the  cinders  not  be  equal  to  the  copper,  sufficient 
unroasted  ore  is  added  in  the  mill  to  bring  it  up  to  1.5  times  the  copper 

Then  9600  lb.  of  mixture  is  taken  as  a  furnace  charge,  and  is  heated 
to  a  very  dull-red  heat,  say  800°  F.,  and  well  stirred.  When  properly 
worked  such  a  change  will  be  finished  in  about  8  hours  making  three 
charges  every  24  hours  per  furnace.  The  charge,  when  finished,  is  drawn 
from  the  furnace  to  the  floor  to  cool,  and  then  transported  in  tram  cars  to 
the  leaching  vats. 

MuSle  furnaces  have  almost  entirely  superseded  the  old  reverberatory 
furnaces  for  chloridizing  roasting.  Contrasting  the  action  of  the  two 
types  of  furnaces,  the  muffle  possesses  the  advantage  of  requiring  but  one- 
half  the  condensing  capacity,  as  only  the  gases  resulting  from  the  chlo- 
ridization  pass  through  the  towers  while  in  the  reverberatory  the  gases 
of  combustion  as  well  must  pass  through.  Apart  from  this  decided 
advantage  of  the  muffle  furnace,  better  results  are  obtained,  as  both  the 
gases  from  the  reaction  and  those  from  the  combustion  are  under 
separate  control. 

The  construction  of  the  modern  muffle  furnace  is  well  illustrated  in 
detail  in  Fig.  49.     The  arch  at  the  fire-box  is  made  double  to  prevent  the 
'  Min.  Industry,  Vols.  VIII  and  IX. 






cinders  from  becoming  overheated  at  that  end,  and  the  fire  flue  is  placed 
beneath  the  furnace  rather  than  at  the  end  of  the  passageway,  which 
is  constantly  traversed  by  the  workmen.  A  bridge  wall  is  constructed 
beneath  the  bed  of  the  hearth  of  the  muffle  and  guides  the  heat  directly 
beneath  it  to  the  entrance  of  the  underground  flue  at  the  fire-box  end 
where  the  fire  flue  damper  for  controling  the  draft  is  located.  The 
furnace  has  doors  on  both  sides  of  the  fire  and  muffle.  The  furnace 
shown  in  the  illustration  was  designed  for  the  treatment  of  cinders  con- 
taining less  than  2  per  cent,  copper,  with  a  normal  amount  of  sulphur, 
the  charge  being  of  usual  weight.  The  hoppers  hold  sufficient  mate- 
rial for  an  entire  charge. 

In  the  construction  of  the  tanks  every  supporting  timber,  plank, 
and  pin  should  be  painted  on  all  sides  with  hot  soft  tar  before  being  put 

Plug  Tap 

Fig.  50. — Details  of  tank  construction  for  leaching  cupriferous  pyrites  cinders. 

in  place.  This  is  absorbed  by  ~the  wood  and  protects  it  against  the 
destructive  action  of  the  acid  liquors.  The  tanks  should  be  constructed 
of  3-in.  plank,  with  an  inside  shell  of  the  same  thickness,  and  a  space  of 
3  in.  between  the  plank  and  the  shell.  All  should  be  put  together  with 
wooden  pins  and  bound  together  as  shown  in  the  accompanying  drawing. 
Fig.  50.  The  space  between  the  tanks  and  shells  should  be  filled  with  a 
mixture  of  hard  tar  and  sand.  The  bottom  should  be  covered  to  the 
depth  of  3  in.  with  the  mixture  fused  to  that  between  the  shell  and  tank. 
This  bottom  covering  should  be  protected  against  wear  of  shovels  by 
a  layer  of  chemical  brick  laid  in  cement. 


The  holes  for  drawing  off  the  liquor  are  bored  through  wooden 
blocks  6X6  in.  square  set  inside  of  the  tank  near  the  bottom,  and  pro- 
vided with  wooden  spigots  or  plugs,  extending  to  the  weak  and  strong 
liquor  launders.  All  the  launders  are  made  preferably  of  very  sappy 
yellow  pine,  dug  out,  with  the  ends  or  joints  halved  together  and  caulked 
with  oakum  and  red  lead. 

The  proper  dimensions  for  the  leaching,  settling,  and  precipitating 
tanks  are,  12X12X4  ft.;  12x12X6  ft.,  and  12X12X6  ft.,  respectively. 

In  preparing  the  mixture  of  hard  tar  and  sand,  composed  of  about 
equal  parts,  care  must  be  exercised  not  only  to  produce  a  homogeneous 
mass,  but  also  that  all  the  moisture  and  air  be  expelled.  The  usua 
method  of  procedure  is  to  heat  the  hard  tar  in  a  large  kettle,  and  stir  it 
until  the  moisture  and  air  have  been  expelled,  then  adding  the  sand  hot 
and  stirring  the  mixture  until  the  desired  result  is  obtained.  The  sand 
must  of  course  be  first  screened,  and  have  all  combustible  matter  burned 
out  of  it  before  mixing  with  the  tar. 

The  bottom  of  the  leaching  tanks  are  provided  with  hard-burned  red 
brick  laid  flat  side  by  side,  and  covered  with  old  hay  or  small  pieces  of 
refuse  coke.     This  makes  a  quite  durable  filter. 

The  roasted  mixture,  charged  into  the  tanks,  is  first  leached  with  weak 
liquor  from  a  previous  operation,  and  then  with  water  and  dilute  hydro- 
chloric acid  from  the  condensing  towers  connected  with  the  furnaces. 
After  leaching  is  completed  the  residue  (purple  ore  or  blue  billy)  is 
shoveled  directly  into  gondola  cars  for  shipment  to  the  iron  smelters. 

All  the  copper  liquors  are  run  into  settling  and  storage  tanks;  the 
weak  liquor  being  pumped  or  blown  back  with  steam  injectors,  to  the 
leaching  tanks  when  required.  Any  liquors  of  18°  B.  and  upward  are 
left  in  the  settling  tanks  until  all  the  lead  sulphate,  etc.,  has  settled  out. 
The  lead  sulphate  usually  contains  some  gold;  indeed  an  average  of 
SlOO  per  month  is  not  an  unusual  recovery  in  a  plant  of  the  capacity 
described,  and  it  is  therefore  essential  for  the  recovery  of  this  as  well  as 
for  other  reasons  that  everything  that  will  settle  out  of  the  liquor,  be 
given  time  to  settle  in  three  tanks.  A  part  of  these  tanks  may  also  be 
used  for  precipitating  the  silver  with  the  iodide,  in  Claudit's  process,  but 
since  in  practice  this  process  leaves  an  average  of  5  oz.  silver  in  a  ton  of 
precipitated  copper,  and  since  electrolytic  copper  works,  as  well  as  blue 
vitriol  works,  handling  or  using  silver-  and  gold-bearing  copper,  will 
pay  for  at  least  95  per  cent,  of  the  silver  content  and  full  market  value 
of  the  copper  and  gold  contents  of  the  cement  copper,  this  part  of  the 
copper  extraction  does  not  pay,  especially  since  the  cost  per  annum  of 
precipitating  the  silver  by  the  iodide  method  in  a  plant  of  this  size  is  not 
less  than  $12,000. 

After  the  lead  sulphate,  etc.  has  settled  out  of  the  strong  liquors, 
they  are  run  into  the  copper  tanks  and  the  copper,  silver,  and  gold  con- 



tents  are  precipitated  by  means  of  clean  thin  scrap  iron  from  the  rolling 
mills.  These  tanks  should,  have  wooden  slats  so  placed  as  to  form  an 
open  false  bottom  about  two  feet  above  the  real  bottom,  for  the  support 
of  the  scrap  iron.  Live  steam  is  let  into  the  liquor  during  the  process 
of  precipitation.  The  steam  serves  to  accelerate  the  process,  and  also 
keeps  the  liquor  in  motion,  and  washes  off  the  copper  from  the  iron  as 
fast  as  it  is  precipitated.  The  copper  naturally  finds  its  way  between 
the  slats  to  the  bottom  of  the  tank,  and  when  it  is  desirable  to  remove 
it,  the  remaining  scrap  iron  can  readily  be  removed  from  the  false  bottom 
practically  free  from  copper.  The  cement  copper,  on  the  bottom  of  the 
tank,  after  washing  through  perforated  cast-iron  plates  set  in  a  frame 
over  the  tanks,  assays  about  90  per  cent.  Cu;  35  oz.  Ag.  and  0.15  oz.  Au, 
per  ton.  The  precipitated  chloride  solutions  are  then  run  into  the 
sewer,  by  first  passing  it  through  a  series  of  tanks  in  the  ground, 
which  are  filled  with  scrap  iron,  to  recover  any  copper  and  silver 
which  may  have  been  left  in  the  waste  chloride  solutions  and  the  wash 

The  cost  of  treating  Spanish  pyrites  cinders,  as  above  described, 
at  Natrona,  Pa.,  in  a  works  having  a  daily  capacity  of  say  200  tons  of 
mixture  (cinder  and  salt)  will  be  about  as  follows: 

2  samplers, 

at  $2 . 50 


6  mill  men, 




1  mechanic, 




2  engineers, 




3  firemen, 




4  weightmen  and  furnace  chargers, 




28  furnace  men, 




1  hoistman, 




27    furnace    material    handlers,    coal,    and 






2  leachers. 




4  copper  precipitators, 




Unloading  cinders  and  salt, 


Loading  purple  ore  for  shipment. 


21  tons  of  salt, 




Pyrites  fines. 


20  tons  of  coal. 




5  1/2  tons  of  sheet-iron  scrap, 




Repairs,  depreciation,  management,  etc. 



Or  $1.87  per  ton  of   2000  lb.    pyritic   cinders,    or  $1.69   per  ton  of 
2000  lb.  of  mixture. 

Cost  of  Producing  Copper  by  the  Longmaid-Henderson  Process  in  a 
Modern  Plant,  Using  Mechanical  Roasters.— The  cost,  per  pound  of 
copper   extracted,   in   a   modern    Longmaid-Henderson   process   plant 


located  in  the  Eastern  United  States,  using  mechanical  furnaces,  and 
treating  from  300  to  400  tons  of  pyritic  cinders  per  day  having  a  copper 
content  of  2.27  per  cent.,  and  a  sulphur  content  of  2.28  per  cent.,  is  as 

Cost,  Per  Pound  of  Copper  Extracted  in  Modern  Longmaid-Henderson 


Items  of  expense, 

Per  pound  of  copper 

Process  labor, 


Misc.  supplies, 




Fuel  oil. 




Scrap  iron, 


Repairs,  all  labor 

and  material; 


Total,  $0.0519 

The  tailings  are  worked  up  by  agglomeration  into  an  iron  ore  of  the 
following  composition: 

Metallic  iron,  68 .  00  per  cent. 

Sulphur,  0 .  07  per  cent. 

Copper,  0. 15  per  cent. 

Silica 3 .  00  per  cent. 

Phosphorous,  0.012  per  cent. 

Oxygen,  etc.,  28.77  per  cent. 

In  the  Western  United  States,  under  somewhat  similar  conditions, 
in  treating  material  having  a  copper  content  of  2.27  per  cent.,  the  esti- 
mated cost  per  ton  of  2000  lb.  is  as  follows: 

Items  of  expense, 

Process  labor, 

Misc.  material, 


Lubricating  oil. 

Fuel  oil, 


Scrap  iron, 

Repairs,  labor  and  material, 

Total,  $2.96 

Elimination  of  Arsenic,  Antimony  and  Bismuth.* — In  the  iongmaid- 
Henderson  process  the  chloridization  of  the  copper  preparatory  to  its 
solution,  is  accompanied  by  the  three  elements,  arsenic,  antimony,  and 
bismuth.  These  chlorides  being  volatile  at  low  temperatures  are  carried 
off  to  a  greater  or  less  extent,  and  are  largely  dissolved  in  the  wash  water 
and  collected  with  the  condensed  acids,  in  the  tower  liquors.     The  arsenic 

'Trans.  A.  I.  M.  E.,  Vol.  XXXIII,  p.  667,  Allen  Gibb. 


per  ton  of  ore 



















in  a  notable  quantity  and  the  antimony  in  a  less  amount  mainly  in  com- 
bination as  arsenates  and  antimonates,  with  some  bismuth  remain  m  the 
roasted  ore.  In  washing  with  water,  these  salts,  as  well  as  the  bismuth 
that  remains  in  the  roasted  ore,  are  dissolved  only  in  minute  traces,  so  that 
if  the  copper  is  precipitated  from  these  solutions  it  would  be  practically 
free  from  impurities.  The  copper,  insoluble  in  water,  that  is  in  the 
roasted  ore,  is  removed  by  the  use  of  tower  liquors,  and  under  the  action 
of  this  solvent  a  considerable  proportion  of  the  arsenic,  antimony,  and 
bismuth  that  remain  in  the  calcined  ore  is  dissolved.  This  increases  the 
proportion  of  these  elements  that  is  already  present  in  the  tower  liquors. 

In  the  tower  liquors,  obtained  by  washing  the  gases  from  the  fur- 
naces, there  was  found:  arsenic,  0.0222  gm.  antimony,  0.0005  gm.;  and 
bismuth,  0.0046  gm.  per  liter. 

The  proportion  of  impurities  in  the  copper  from  the  same  ore  vary 
greatly  according  to  whether  the  practice  of  treating  the  aqueous  and 
acid  solutions  separately  is  followed  or  not. 

The  following  data  is  taken  from  practice  in  which  the  two  solutions 
were  not  kept  separate: 

Roasted  ore                                           Precipitate 

Per  cent, 

Per  cent, 
Cu  =  100  per 

Per  cent, 

Per  cent. 


Cu  =  100  per 


Total  per  cent, 
of  elimination 



Antimony. . .  . 






Extraction  of  Copper  from  Atacamite.' — At  Chiquicamata,  Chile,  a 
multitude  of  small  fissures,  filled  with  atacamite,  or  oxychloride  of  copper, 
traverse  the  country  rock,  consisting  of  granites,  pegmatites,  syenites,  in 
every  direction.  This  kind  of  deposit,  which  appears  to  be  in  the  nature 
of  a  stock  work,  a  leaching  process  is  usually  adopted  in  order  to  extract 
the  metal  from  the  oxychloride. 

In  treating  Atacamite  ores  the  difficulty  has  been  in  the  filtration  of 
the  liquid  containing  the  dissolved  copper.  The  chemical  reactions 
which  take  place  between  the  perchloride  and  oxide  of  iron  on  the  one 
part  and  the  ferric  solution  and  argilaceous  portion  of  the  gangue  on  the 
other,  result  in  the  formation  of  a  gelatinous  precipitate,  which  has  to  be 
washed  many  times  in  order  to  get  out  the  dissolved  metal. 

^London  Mining  Journal,  June  30,  1906,  Nioanor  Argandona. 


The  first  part  in  the  treatment  of  the  ore  by  a  new  process  consists  in 
converting,  by  the  action  of  steam,  a  portion  of  the  chloride  of  copper  in 
the  ore  into  hydrochloric  acid  and  black  oxide  of  copper,  according  to 
the  equation: 

CuCl2  +  H20  =  CuO+2HCl. 

A  certain  portion  of  the  ore  is  put  into  large  clay  retorts,  or  into 
iron  retorts  lined  with  a  thin  coating  of  clay,  and  exposed  at  a  temper- 
ature of  230°  C,  to  the  action  of  steam.  Theoretically,  there  is  sufficient 
water  of  combination  in  the  ore  itself  to  supply  the  quantity  of  vapor 
needed.  The  expenditure  of  steam  is  very  small,  but  a  certain  quantity 
is  necessary  in  order  to  accelerate  the  reaction.  The  black  oxide  result- 
ing from  this  process  is  reduced  by  smelting. 

The  second  part  of  the  process  consists  in  submitting  the  ore,  which 
contains  from  3  to  4  per  cent,  copper,  to  the  action  of  hydrochloric  acid 
obtained  as  above  described.  For  this  purpose  cylindrical  wooden  or 
brick  vats  are  used,  having  special  lining  of  pitch.  Ordinary  filters  can 
be  employed  in  this  operation,  as  no  gelatinous  precipitate  is  produced. 


Iron.— Iron  is  almost  universally  used  in  the  chemical  precipitation  of 
copper  from  its  solutions.  It  presents  many  advantages  over  other  pre- 
cipitants;  it  is  usually  quite  cheap,  generally  obtainable,  and  gives  the 
resultant  copper  in  the  metallic  condition.  Scrap  iron  is  ordinarily  em- 
ployed, but  scrap  iron,  while  cheap  enough  in  industrial  centers  becomes 
prohibitive  in  mining  districts  which  are  located  long  distances  from 
railroads  and  from  the  source  of  iron  supply. 

For  the  precipitation  of  copper  by  iron,  the  solution  should  be  as  free 
as  possible  from  acid  or  ferric  salts.  The  acid  and  ferric  salts  act  on  the 
iron  and  waste  it  without  precipitating  any  copper.  Theoretically  88.8 
lb.  of  iron  are  required  by  weight  to  precipitate  100  lb.  of  copper  from 
sulphate  solutions.  In  practice,  the  consumption  greatly  exceeds  the 
theoretical  amount,  because  the  solution  cannot  be  kept  free  from  acid 
or  ferric  salts  and  because  scrap  iron  which  is  contaminated  more  or 
less  with  impurities,  not  available  for  precipitation  is  ordinarily  used. 
With  care,  the  consumption  of  iron  in  precipitating  copper  from  sulphate 
solutions,  in  practice,  will  average  about  1.5  lb.  of  iron  per  pound  of 
copper  precipitated;  under  the  adverse  conditions  of  free  acid,  ferric 
salts,  and  impure  scrap  iron,  the  consumption  of  iron  may  rise  to  2  or 
even  3  parts  of  iron  to  1  part  of  copper. 

Ferrous  sulphate  by  prolonged  contact  with  air  is  decomposed  into 
free  sulphuric  acid  and  ferric  sulphate;  the  former  dissolves  iron,  and  the 
latter  combines  with  it  to  again  form  ferrous  sulphate.  The  following 
equations  explain  the  principal  chemical  changes  that  take  place,  in 
precipitating  with  iron  from  a  sulphate  solution. 

(1)     CuS0,+Fe  =  Cu+FeS04. 

This  represents  the  theoretical  reaction,  without  any  interfering 
elements,  and  if  it  could  be  theoretically  carried  out,  only  88.8  parts 
of  iron  would  be  required  to  precipitate  100  parts  of  copper.  As  both 
ferric  sulphate  and  free  sulphuric  acid  are  likely  to  be  in  the  solutions, 
the  following  reactions  also  take  place: 

(2)  Fe2(SOj3+Fe  =  3FeSO,. 

(3)  H,S0,+Fe=FeS0,  +  2H. 

The  excess  of  iron  consumed  may  be  said  in  general  to  be  due  to  the 
quantity  of  ferric  iron  and  free  acid  in  the  copper  solution.     The  tendency 



of  the  iron  is  to  reduce  the  ferric  to  the  ferrous  salts,  when  the  theoret- 
ical amount  will  be  more  nearly  approached. 

Some  of  the  ferric  sulphate  may  not  be  directly  reduced  by  the  iron, 
but  may  act  on  the  precipitated  copper,  as  shown  by  the  following 

Fe^  (SO ,)  3  +  Cu  =  CuSO ,  +  2FeS0  „ 

but  as  the  copper  so  dissolved  has  to  be  precipitated  at  the  expense  of 
the  iron,  the  ultimate  amount  of  iron  consumed  in  reducing  the  ferric 
salts  is  the  same. 

An  accurate  determination  of  the  iron  consumed  per  ton  of  copper 
produced  can  be  arrived  at  by  testing  the  ingoing  and  outgoing  solutions 
to  the  precipitators,  to  determine  the  amount  of  the  copper  precipitated, 
of  the  ferric  iron  reduced,  and  of  free  sulphuric  acid  neutralized,  and 
calculating  the  quantity  of  metallic  iron  required  to  bring  about  these 
changes.  S.  R.  Adcock '  gives  the  following  partial  analysis  and  calcu- 
lations, of  samples  of  liquor  entering  and  leaving  one  of  the  cementation 
tanks  at  Rio  Tinto. 



'  Entering  ''  Leaving 

Copper I  2.064  j  3 

Ferriciron i  1.328  Nil 

Sulphuric  acid j  1 .  198  |  712 

Calculations  in  Grams  per  Cubic  Meter  of  Liquor 
Cu  precipitated,  2.064-3=2.061x8/9  =  1.832  grm.  of  iron  required. 

Ferric  iron  reduced,  1.328  X  1/2  =    644  grm.  of  iron  required. 

Sulphuric  acid  neutralized,  1.189-712  =  486X4/7=    278  grm.  of  iron  required. 
Total  iron,  2^74  grm. 

2.061  grm.  of  copper  precipitated  would  require  2.774  grm.  of  iron 
or  one  part  of  copper  precipitated  from  the  liquor  would  require  1.345 
parts  of  iron.  Taking  the  metallic  contents  of  the  pig  iron  used  at  92 
per  cent,  the  consumption  in  this  instance  works  out  at  1  part  of  copper  to 
1.462  parts  of  pig  iron.  In  addition,  there  is  a  small  consumption  of 
iron,  due  to  impurities  in  the  solution  and  waste  in  cleaning  up  the 

It  is  desirable,  for  the  best  work  in  precipitation,  to  have  the  copper 
solution  slightly  acid.  A  slight  acidity  tends  to  hasten  the  precipitation, 
and  prevents  the  separation  of  basic  iron  salts.     In  general,  the  acid  con- 

'  Min.  Ind.,  Vol.  IX,  p.  238. 


tent  of  the  copper  solution  for  precipitation  in  a  running  stream,  or  if 
agitated,  should  not  be  more  than  0.1  to  0.2  per  cent.  If  the  acid  much 
exceeds  this  amount  the  consumption  of  iron  is  likely  to  be  correspond- 
ingly high. 

Every  precaution  should  be  taken  to  reduce  the  ferric  iron  to  the 
ferrous  condition  before  precipitating  the  copper,  for  in  most  instances 
the  excessive  consumption  of  iron  during  precipitation,  can  be  traced  to 
the  high  ferric  iron  content  of  the  liquor  under  treatment. 

If  there  is  arsenic  in  the  ore  it  is  likely  to  go  into  solution  with  the 
copper,  and  is  to  a  certain  extent  precipitated  with  it.  When  the  liquor 
is  rich  in  copper  and  the  precipitation  is  taking  place  rapidly,  the  amount 
of  arsenic  precipitated  is  comparatively  small,  but  as  the  liquor  gets 
weaker  in  copper,  the  proportion  of  arsenic  to  copper  precipitated  is 
much  higher. 

If  the  copper  is  dissolved  as  a  chloride,  either  by  hydrochloric  acid  or 
metal  chloride,  the  precipitation  may  take  place  either  from  the  cupric 
or  cuprous  chloride,  as  shown  by  the  following  reactions: 

CuCl^  +Fe=reCl2  +  Cu. 
Cu,Cl2+Fe=reCl2  +  2Cu. 

It  is  evident,  therefore,  from  these  equations,  that  twice  as  much  copper 
is  precipitated,  theoretically,  per  unit  of  iron  from  cuprous  chloride  as 
from  cupric  chloride.  The  precipitation  from  the  cupric  chloride  is 
theoretically  the  same  as  that  of  cupric  sulphate — 88.8  parts  of  iron  are 
required  to  precipitate  100  parts  of  copper — whereas  only  44.4  parts 
of  iron  are  required  to  precipitate  100  parts  of  copper  from  cuprous 

In  practice,  at  Stadtberg,  127  parts  by  weight  of  iron  were  required 
to  precipitate  100  parts  of  copper  from  a  chloride  solution  in  which 
hydrochloric  acid  was  used  as  the  solvent.  In  the  Hunt  and  Douglas 
process,  in  which  the  copper  is  precipitated  from  cuprous  chloride,  ohly 
from  50  to  70  parts  of  iron  are  required  to  precipitate  100  parts  of  copper. 

If  there  is  free  hydrochloric  acid  or  ferric  chloride  in  the  solution,  iron 
will  combine  with  them  to  form  ferrous  chloride,  and  will  consequently 
be  wasted,  as  in  the  analogous  procedure  in  the  sulphate  solution;  this 
may  be  shown  by  the  following  equations: 

2HCl+Fe=FeCl,  +  2H. 
2FeCl3+Fe  =  3FeCl2. 

The  neutralization  of  the  free  acid,  and  the  reduction  of  the  ferric 
to  the  ferrous  chloride,  is  quiet  as  necessary  as  with  sulphate  solutions  to 
effect  the  best  economy  in  the  precipitation. 

The  iron  for  precipitation,  whether  from  a  sulphate  or  chloride  solu- 
tion may  be  used  in  the  form  of  wrought,  pig  iron,  iron  bars,  or  iron  spong. 


Pulveront  iron  in  the  form  of  ground  sponge  acts  most  rapidly.  Bar 
iron  yields  a  cgarse  grained  cement  copper,  with  but  little  coherence; 
gray  pig  iron,  which  acts  faster  than  white  iron,  gives  a  more  pulverent 
precipitate,  while  white  iron  throws  down  coherent  masses.  The 
graphite  in  the  pig  iron  separates  out  during  the  precipitation  and  renders 
the  cement  copper  impure. 

Sponge  Iron. — It  is  evident  that  if  iron  is  used  as  the  precipitant  of 
copper,  it  would  be  quite  an  advantage  if  the  iron  could  in  some  way  be 
cheaply  manufactured  at  the  copper  reduction  works.  The  purchase  of 
scrap  or  pig  iron,  and  its  transportation  to  the  reduction  works  at  the 
mines,  is  usually  a  prohibitive  expense.  An  ordinary  iron  smelting 
plant  could  not  be  considered. 

The  manufacture  of  sponge  iron,  in  which  a  high-grade  ore  is  reduced 
to  metal  without  the  necessity  of  smelting  or  fluxing,  has  in  a  measure 
solved  the  problem,  but  notwithstanding  that  the  manufacture  of  sponge 
iron  and  its  application  to  copper  precipitation  has  long  been  known,  its 
use  has  not  met  with  much  encouragement. 

The  principle  of  the  manufacture  of  sponge  iron  is  quite  simple.  If 
ferric  oxide,'  FcjOg,  is  heated  in  a  highly  reducing  atmosphere,  the 
oxygen  of  the  iron  oxide  combines  with  the  reducing  gases,  and  the 
resultant  product  is  finely  divided  metallic  iron,  containing  more  or  less 
impurities.  Great  care  must  be  used  in  cooling  the  iron  to  prevent 
reoxidation,  so  that  the  cooling,  as  well  as  the  heating  must  be  done  in 
the  presence  of  reducing  gases. 

The  following  description  by  Lunge*  gives  the  details  of  early  European 
practice  and  its  application  to  copper  precipitation. 

"The  precipitation  of  copper  from  its  solutions  by  means  of  iron  takes  place 
more  rapidly  by  employing  spongy  iron,  as  was  done  at  the  Beds  Metal  Works. 
This  product  is  made  by  reducing  ferric  oxide  at  so  low  a  temperature  that  the 
iron  cannot  combine  with  the  carbon  and  cannot  melt,  but  remains  in  the  finely 
divided  state  as  'sponge.'  This  method  was  tried  in  England  for  the  first  time 
in  1837.  Gossage,  in  1859,  was  the  first  to  use  it  in  the  wet  method  of  coppfr 

"The  furnace  usually  employed  in  making  sponge  iron,  is  a  reverberatory  in 
which  the  flame,  after  having  passed  directly  over  the  charge,  returns  below  the 
furnace  bed,  and  thus  heats  the  charge  indirectly  from  below.  Figs.  .51  and  .52 
show  the  essential  details  of  the  furnace.  It  is,  in  the  drawing,  28  ft.  9  in.  long; 
the  working  bed  has  a  length  of  22  ft.  and  a  width  of  8  ft.  Dwarf  walls,  a  a,  9  in. 
high,  divide  it  into  three  compartments,  which  on  one  .side  have  two  working 
doors,  6  b,  each.  Each  compartment  is  charged  and  finished  by  itself.  The 
working  doors  are  of  cast  metal  and  run  in  air-tight  frames ;  the  same  is  the  case 
with  the  fire  doors.  The  fireplace  is  constructed  for  generating  a  reducing  flame ; 
the  grate  has  a  surface  4  X  3  ft. ;  and  the  bearers  d,  are  3  ft.,  latterly  even  4  ft.  8  in. 

'  "Sulphuric  .Vcid  and  .\lkali  Manufacture,"  p.  815. 






below  the  fire  bridge ;  so  that  a  very  deep  layer  of  fuel  is  obtained,  which  does 
not  allow  any  oxygen  to  get  inside  the  furnace.  The  furnace  bed  is  formed  by 
fire  tiles  4  in.  thick,  with  rabbited  edges,  partly  resting  upon  the  walls  forming 
the  divisions  of  the  lower  flues,  partly  upon  railway-bars. 

"The  flame  having  traveled  through  these  flues,  descends  in  a  vertical  shaft 
along  the  fire  bridge,  and  thence  goes  to  the  chimney.  In  this  descending  shaft 
there  is  a  fireclay  damper,  which  is  closed  every  time  before  a  working  door  or 
fire  door  is  opened.  The  9-in.  furnace  roof  is  surmounted  by  a  flat  cast-iron 
dish,  supported  by  short  pillars,  for  drying  the  ore  and  mixing  it  with  coal; 
the  mixture  is  charged  into  the  furnace  through  the  6-in.  pipes,/,  carried  through 

Fig.  52. — Sponge-iron  furnace.     Transverse  section. 

the  arch.  The  whole  furnace  rests  on  brick  pillars,  g,  and  the  floor  on  the  working 
sides  must  be  so  much  higher  than  that  on  the  discharge  side,  that  the  discharging 
boxes  can  be  run  underneath  the  furnace  between  the  brick  pillars.  The  dis- 
charging takes  place  through  6-in.  pipes,  h,  descending  in  front  of  the  working 
doors  through  the  furnace  bottom  andthe  lower  flues. 

"The  discharge  boxes  are  made  of  sheet  iron,  of  rectangular  section,  tapering 
toward  the  top.  The  cover  is  fast,  and  has  in  its  center  a  6-in.  opening  with 
upright  flange,  by  which  the  box  is  connected  with  the  discharge  tube.  The 
bottom  of  the  box  is  movable,  and  turns  on  one  side  on  hinges,  while  the  other 
side  is  fastened  by  bolts  and  cotters.  The  whole  is  mounted  on  four  wheels  in 
such  a  way  that  they  do  not  interfere  with  the  movements  of  the  bottom.  Each 
box  has  a  capacity  of  12  cu.  ft. 

"  When  the  furnace  is  at  a  bright  red  heat,  it  can  be  charged.  Y.Avh.  compart- 
ment receives  a  charge  of  2000  lb.  of  'purple  ore'  and  600  lb.  of  coal,  which  have 
passed  through  a  sieve  with  eight  holes  to  the  lineal  inch. 

"The  charging  takes  place  from  the  cast-iron  dish  above  the  furnace  roof. 


The  fire  and  working  doors  are  closed,  so  that  the  air  enters  solely  through  the 
coals  on  the  grate,  care  being  taken  that  the  burning  mass  does  not  become 
hollow,  lest  uncombined  oxygen  should  get  inside  the  furnace.  The  time  of 
reduction  in  the  compartment  nearest  the  fire  bridge  varies  from  9  to  12  hours; 
in  the  second  it  is  about  18  hours,  and  in  the  third  about  24  hours.  The  depth 
of  the  charge  lying  on  the  bed  is  about  6  in.  During  the  time  of  reduction  each 
compartment  must  be  turned  over  twice,  or  even  three  times.  Although  during 
this  time  the  damper  is  closed,  a  little  air  always  enters  the  furnace ;  but  the  turn- 
ing over  is  indispensable,  as  the  mass  would  otherwise  cake  together.  The  time 
above  stated  refers  to  a  bright  red  heat;  a  low  heat  is  sufficient  for  reduction  and 
the  iron  thus  made  is  even  better  for  the  precipitation  of  copper;  but  as  in  this 
case  much  more  time  is  required  for  reduction  (up  to  60  hours)  this  style  of 
working  does  not  pay.  The  fire  place  being  so  deep,  fresh  coal  need  only  be 
thrown  on  twice  or  three  times  every  12  hours,  say  1500  lb.  per  ton  of  ore. 

"The  completion  of  the  reduction  is  ascertained  by  testing.  A  small  sample 
is  taken  out,  put  on  an  iron  plate,  covered  with  a  brick  until  it  has  become  cold, 
and  1  grm.  of  the  (unoxidized)  central  part  tested  by  a  cupric  sulphate  solution 
of  known  strength,  which  is  run  from  a  burette  on  the  spongy,  iron  with  frequent 
stirring;  from  time  to  time  a  drop  is  put  on  a  bright  blade  of  iron,  to  see  whether 
any  stain  of  copper  is  produced  upon  it.  When  the  reaction  in  any  of  the  three 
compartments  is  finished,  the  damper  is  closed;  two  of  the  discharging  boxes  are 
run  underneath  the  furnace,  and  their  openings  connected  with  the  discharge 
pipes  by  an  iron  hoop  luted  with  clay;  then  the  charge  is  raked  down  into  the 
boxes  as  quickly  as  possible.  The  boxes  are  then  closed  with  the  loose  cover, 
run  out  again,  and  allowed  to  cool  for  48  hours.  They  are  then  lifted  by  a  crane, 
and  the  cotters  are  knocked;  whereupon  the  bottom  turns  on  its  hinges  and  the 
whole  mass  of  spongy  iron  readily  falls  out,  owing  to  the  box  tapering  upward. 
The  sponge  is  then  finely  ground  in  a  heavy  chillean  mill  6  ft.  in  diameter,  and 
passed  through  a  sieve  with  fifty  holes  per  lineal  inch.  It  is  now  ready  for  the 
precipitation  of  the  copper. 

The  following  is  an  analysis  of  sponge  iron  produced  from  purple  ore,  from 
Spanish  pyrites. 

Purple  Ore  Sponge  Iron 

Ferric  oxide,  95 .  10  per  cent.  8 .  15  per  cent. 

Ferrous  oxide,  2.40  per  cent. 

Metallic  iron,  70 .  40  per  cent. 

Copper,  0.18  per  cent.  0 .  24  per  cent. 

Lead  oxide,  0.96  per  cent 

Lead,  0.27  per  cent. 

Sulphur,  0.07  per  cent.  0  07  per  cent. 

Calcium  oxide,  0.20  per  cent. 

Sodium  oxide,  0. 13  per  cent.  

Sulphur  trioxide,  0 .  78  per  cent 

Alumina,  .  :  .  . 0 .  19  per  cent. 

Zinc,  0.30  per  cent. 

Silicious  residue,  2 .  13  per  cent.        9 .  00  per  cent. 

99.55  per  cent.      99.62  per  cent. 


"  When  spongy  iron  is  employed  for  the  precipitation  of  copper,  cont'nuous 
stirring  is  required,  for  which  at  some  works  a  mechanical  agitator  is  used,  at 
others,  manual  labor." 

Any  good  iron  ore  may  be  converted  into  iron  sponge,  or  the  con- 
centrate residues,  after  roasting  and  copper  extraction. 

Crucible  Method  of  Iron  Ore  Reduction. — Another  method  of  reducing 
iron  ore  in  connection  with  copper  precipitation — that  in  which  the 
mixture  of  ore  and  coal  is  reduced  in  a  crucible — might  be  considered. 
This  method  of  manufacturing  iron  has  been  in  use  in  China  from  remote 
antiquity,  and  large  quantities  are  in  this  way  reduced  there. 

At  Shansi,  China,  the  crucibles  are  about  19  in.  high  and  6.5  in.  in 
diameter..  They  are  filled  with  a  mixture  of  ore  which  has  been  broken 
and  sorted  to  walnut  size;  coal  of  about  the  same  size,  and  coal  dirt. 
The  proportions  are  four  baskets  of  ore,  one  basket  of  coal,  and  one 
basket  of  coal  dirt  on  top.  The  crucibles  are  heated  in  a  stall  furnace. 
It  takes  about  16  hours  to  reduce  the  iron.  After  cooling  the  crucibles 
are  removed  and  broken  and  the  iron  taken  out. 

Precipitation  with  a  Coke-iron  Couple. — ^^'.  L.  Austin'  gives  the  result 
of  a  30-day  comparative  test  on  a  working  scale,  using  ordinary  mine 
waters,  between  iron  and  the  coke-iron  couple  as  the  precipitant.  The 
average  daily  assays  showed  the  following  extraction  effected  within 
8  ft.  from  where  the  waters  entered  the  precipitation  boxes: 

Percentage  of  copper  removed  in  the  boxes  containing  clean  wrought 
iron 25 

Percentage  of  copper  removed  in  the  boxes  containing  coke-iron 
couple 43 

It  was  found  that  when  coke  was  added  to  the  iron  in  the  precipitating 
boxes  through  which  the  sulphate  mine  waters  were  flowing,  by  actual 
weight  from  1.15  to  2.66  lb.  copper  was  precipitated  to  the  pound  of  iron 
consumed.  One  of  the  advantages  in  employing  the  coke-iron  couple 
in  cementation  is  that  the  precipitation  of  the  copper  takes  place  within 
a  short  distance  from  the  point  where  the  liquors  enter  the  boxes,  thereby 
avoiding  to  a  great  extent  the  formation  of  basic  iron  salts  which  usually 
debase  the  cement  copper. 

Precipitation  with  a  Copper -iron  Couple. — If  iron  plates  or  a  bundle 
of  scrap  iron  in  a  crate  are  immersed  in  a  copper  sulphate  solution  and 
connected  electrically  with  a  copper  plate  to  serve  as  a  cathode,  no  copper 
is  precipitated  on  the  iron,  but  all  is  precipitated  on  the  copper  plate. 
The  iron  acts  as  a  soluble  anode,  going  into  solution  as  ferrous  sulphate, 
while  the  copper  is  deposited  in  a  dense,  practically  pure  condition  on  the 
copper  sheet.  In  this  case  the  iron  and  copper  act  as  a  battery,  with  a 
theoretical  difference  of  potential  of  0.81  volt.     If  more  rapid  precipita- 

'  "Mines  and  Methods,"  January,  1911. 


tion  is  desired  than  given  by  this  action  alone,  a  current  from  a  dynamo 
may  also  be  used  in  connection  with  that  produced  by  the  iron  and  copper. 

Hydrogen  Sulphide. — If  hydrogen  sulphide  is  applied  to  a  copper 
sulphate  solution,  the  copper  is  precipitated  as  cupric  sulphide,  and  an 
amount  of  acid  regenerated  equal  to  that  combined  with  the  copper 
CuS04  +  H3S  =  CuS  +  H2SO,. 

If  the  solution  is  a  chloride,  a  similar  reaction  takes  place: 
CuCl^  +  H^S  =  CuS  +  2HC1. 

In  either  case,  the  regenerated  acid  solution  is  returned  to  the  ore  to 
dissolve  more  copper.  The  precipitated  copper  sulphide,  like  copper 
matte,  may  be  brought  to  blister  copper  in  converters.  The  escaping 
sulphur  dioxide  may  be  used  in  other  stages  of  the  process,  either  in  the 
manufacture  of  sulphuric  acid,  or  in  the  preparation  of  hydrogen  sulphide. 

Many  wet  methods  of  extracting  copper  from  its  ores  have  been  based 
on  the  use  of  hydrogen  sulphide  as  the  precipitant.  Most  of  these  methods 
have  as  their  fundamental  idea,  the  inexpensive  generation  of  the  hydro- 
gen sulphide,  rather  than  a  direct  method  of  copper  extraction. 

At  the  Bede  Metal  Works  the  hydrogen  sulphide  was  produced  by 
treating  sulphide  of  sodium  with  carbon  dioxide,  generated  by  burning 
coke  in  a  shaft  furnace.  The  sulphide  of  sodium  was  made  from  the  acid 
mother  liquor  from  which  the  sulphide  of  copper  had  been  precipitated, 
by  evaporating  to  dryness  in  a  reverberatory  furnace,  mixing  the  residue 
with  coal  dust  and  reducing  it  in  a  similar  furnace.  This  product,  con- 
sisting of  sulphide  and  carbonate  of  soda,  was  leached,  and  the  solution 
treated  with  carbon  dioxide.  The  carbonate  of  soda  was  produced  as  a 
by-product  of  the  process. 

Hydrogen  sulphide,  according  to  Schnabel,'  is  best  generated  by  lead- 
ing sulphur  dioxide  and  water  vapor  over  red  hot  coke  or  charcoal.  For 
this  purpose  the  gases  produced  in  roasting  sulphides  in  pyrite  burners 
or  kilns  are  aspirated  by  means  of  Korting  injectors,  and  forced  together 
with  the  steam  from  the  injectors  through  a  shaft  furnace  filled  with  glow- 
ing charcoal  or  coke.  The  sulphur  dioxide  is  reduced  to  sulphur  by  the 
carbon;  the  water  vapor  forms  with  the  red  hot  carbon,  hydrogen 
and  carbon  monoxide,  and  the  hydrogen  and  sulphur  combine  to  form 
hydrogen  sulphide.  The  coke  or  charcoal  is  kept  hot  by  injecting  a 
stream  of  air  from  time  to  time  as  in  the  production  of  water  gas. 

A  method  devised  by  Sinding  depends  upon  the  action  of  sulphur 
vapor  on  hydrocarbons  and  hydrogen;  another  on  the  decomposition  of 
sodium  sulphide  by  carbon  dioxide.  Sinding  generates  producer  gas  with 
raw  fuel,  and  leads  it  over  red  hot  pyrites;  by  the  action  of  the  hydro- 
carbons and  the  hydrogen  contained  in  producer  gas  upon  the  sulphur 
evolved  from  the  pyrites,  hydrogen  sulphide  is  formed,  and  is  made  to 

^  Handbook  of  Metallurgy,  Vol.  I,  p.  209. 


traverse  a  chamber  in  which  the  cupriferous  solution  is  dropping  down  in 
the  form  of  rain. 

Gill  and  Gelstharp  produced  hydrogen  sulphide  by  the  action  of 
carbon  dioxide  on  sodium  sulphide. 

In  the  application  of  hydrogen  sulphide  to  the  precipitation  of  copper 
solutions,  the  precipitation  is  thorough  and  complete.  If  the  solution 
contains  gold  and  silver,  the  filtrate  will  never  assay  more  than  a  trace; 
that  is  to  say,  less  than  0.0005  oz.  or  1  cent  per  ton.  The  conditions  for 
thorough  precipitation  in  a  normally  working  plant,  are  that  the  solution 
should  be  slightly  acid,  or  if  neutral,  contain  an  excess  of  free  chlorine  or 
ferric  salts.  If  the  solution  is  alkaline  or  neutral,  the  precipitation  of  the 
copper,  gold  and  silver  will  be  complete,  but  some  of  the  undesirable 
elements  will  be  thrown  down  with  them.  The  precipitate  is  voluminous 
and  may  be  difficult  to  filter.  If  the  solution  is  neutral  and  contains  a 
large  excess  of  chlorine  or  ferric  salts,  the  hydrogen  sulphide  reacting 
with  these  substances,  will  reduce  them,  and  acidify  the  solution  suffi- 
ciently to  prevent  the  baser  metals  from  being  precipitated. 

C1  +  H2S=2HC1  +  S, 

2FeCl3  +  H2S  =  2HCl  +  2FeCl2  +  S, 

Fe2(SO,)3-t-H2S=2FeSO,  +  H2SO,  +  S, 

and  thus  hydrogen  sulphide  will  be  consumed,  an  equivalent  of  acid 

regenerated,  and  elemental  sulphur  precipitated  with  the  copper  sulphide. 

The  metals  of  the  alkalies  and  alkaline  earths  are  not  precipitated  from 
either  the  acid  or  alkaline  solutions.  Frequently  compounds  of  calcium 
and  aluminum  will  be  found  in  the  precipitate.  Calcium  sulphate  is 
only  slightly  soluble  in  water,  but  is  somewhat  soluble  in  chloride  solu- 
tions, and  unless  the  solution  is  thoroughly  settled  or  filtered,  it  is  also 
likely  to  be  carried  into  the  precipitate  by  suspension.  If  the  ore  con- 
tains much  alumina  and  the  solution  issuing  from  the  leaching  tanks  is 
neutral  or  alkaline,  a  white  gelatinous  substance,  probably  aluminum 
hydroxide,  settles  out  of  the  clear  or  milky  solution,  and  may  in  this  way 
get  into  the  precipitate.  On  the  addition  of  acid  this  precipitate  is 
redissolved  and  the  solution  becomes  clear. 

Metals  which  are  precipitated  by  hydrogen  sulphide,  as  sulphide,  from 
a  solution  of  their  salts  in  the  presence  of  free  acid,  are : 

Platinum,  color  of  precipitate,  dark  brown. 

Gold,  color  of  precipitate,  dark  brown. 

Silver,  color  of  precipitate,  black. 

Copper,  color  of  precipitate,  black. 

Lead,  color  of  precipitate,  black. 

Tin,  color  of  precipitate,  yellow-brown. 

Antimony,  color  of  precipitate,  orange. 

Arsenic,  color  of  precipitate,  yellow. 

Mercury,  color  of  precipitate,  black. 

Cadmium,  color  of  precipitate,  yellow. 



ifli:t"~"i"           -- 

■■  ■  — ~ 

_.^_ ___._  _ 


Compressed  Air  ^ 


Fig.  53. — Hydrogen  sulphide  generator. 


The  hydrogen  sulphide  employed  in  precipitation  is  usually  generated 
from  iron  sulphide  (FeS),  sulphuric  acid,  and  water.  A  solution  having 
a  large  excess  of  acid  will  require  more  hydrogen  sulphide  to  precipitate 
the  metals  than  one  only  slightly  acid.  A  strongly  acid  solution  will 
give  a  cleaner  precipitate  than  a  neutral  solution,  and  give  less  trouble 
in  the  filter  presses. 

The  hydrogen  sulphide  generator,  for  producing  the  gas  on  a  large 
scale,  from  acid  and  matte,  is  shown  in  Fig.  53  and  may  be  made  of  any 
size  desired.  It  consists  of  an  iron  cylinder  with  flat  cast-iron  bottoms 
and  tops.  If  the  cylinder  is  small,  it  is  made  of  cast  iron;  if  large,  of 
sheet  steel.  It  is  lined  with  lead,  and  all  connections  are  made  of  lead 
pipe  with  burned  joints.  An  equalizing  tank,  also  made  of  lead,  is 
located  close  to  the  generator. 

The  materials  for  generating  the  hydrogen  sulphide  are  charged  in 
the  following  proportions: 

Iron  sulphide  (iron  matte) ,  1     lb. 

Sulphuric  acid,  2.5  lb. 

Water  (approximately),  6     lb. 

The  total  quantity  of  chemicals  charged,  will  depend  on  the  size  of  the 
generator,  and  the  amount  of  metal  to  be  precipitated. 

In  charging  the  generator,  the  exhaust  valve  from  the  equalizing  tank 
is  left  open.  The  ]-equired  amount  of  water  is  then  introduced.  This 
may  be  measured  in  a  small  tank  over  the  generator,  or  it  may  be  run  in 
through  the  water  pipe,  or  hose  inserted  into  the  manhole,  to  a  certain 
depth  which  has  been  determined  before  hand  as  the  right  quantity  of 
water  for  any  required  charge.  The  iron  sulphide,  broken  into  pieces 
about  the  size  of  hens'  eggs,  is  dropped  in  through  the  manhole,  which  is 
then  closed.  The  required  amount  of  acid,  which  had  previously  been 
measured  into  the  small  lead  tank  located  over  the  generator,  is  then  run 
in  and  the  valve  again  immediately  closed.  The  generation  of  hydrogen 
sulphide  begins  at  once.  The  exhaust  valve  on  the  equalizing  tank,  which 
was  left  open  so  that  the  pressure  generated  by  the  gas  would  not  prevent 
the  acid  fr^m  flowing  into  the  generator,  is  then  closed.  The  increasing 
pressure  will  soon  force  the  gas  into  the  copper  precipitating  tanks. 

Iron  sulphide  reacts  with  sulphuric  acid  to  form  hydrogen  sulphide: 

FeS  +  H2SO,=FeSO,  +  H2S. 

The  gas  so  obtained  always  contains  free  hydrogen,  owing  to  the 
presence  of  uncombined  iron  in  the  iron  sulphide. 

Compressed  air  is  used  to  force  the  gas  from  the  generator  into  the 
precipitating  tanks  and  agitate  the  solutions. 

After  the  precipitated  solution  has  settled  long  enough  to  clarify, 
usually  from  4  to  8  hours,  it  is  decanted  through  the  collar  in  the  bottom 


of  the  settling  tank,  into  a  filter  press,  and  the  clear  regenerated  solution 
may  then  be  returned  to  the  ore  to  dissolve  more  copper. 

Lime. — Lime,  calcium  hydrate  or  milk  of  lime,  is  not  suitable  for 
precipitation  from  sulphate  solutions.  It  has  been  used  in  precipitating 
copper  from  cupric  and  cuprous  chloride  solutions,  calcium  chloride  being 

Milk  of  lime  precipitates  cupric  hydrate  from  cupric  chloride,  and 
cuprous  oxide  from  cuprous  chloride.  It  throws  down  copper  as  cupric 
hydrate  from  a  solution  of  the  sulphate,  but  the  precipitate  is  mixed 
with  the  insoluble  calcium  sulphate,  which  is  formed  at  the  same  time, 
and  gives  a  mixture  which  is  very  voluminous  and  very  troublesome  to 
smelt.  The  precipitation  of  the  copper  with  lime  from  cuprous  chloride 
was  for  a  long  time  used  in  connection  with  the  Hunt  and  Douglas 
process,  in  which  the  resulting  calcium  chloride  was  used  in  another 
step  in  the  process,  but  lime  has  nowhere  been  regularly  used  to  pre- 
cipitate copper  from  sulphate  solutions. 


General  Consideration  of    Electrolytic   Methods. 

Definitions. — Electrolysis  may  be  defined  as  the  decomposition  of  a 
chemical  compound  by  the  electric  current.  The  compound  may  be  in 
aqueous  or  igneous  solution. 

Electrolyte  is  a  chemical  compound  in  aqueous  or  igneous  solution 
being  decomposed  by  the  electric  current. 

Electrode  is  a  conductor  to  convey  the  current  of  electricity  into  or  out 
of  the  electrolyte. 

Anode  is  the  electrode  by  means  of  which  the  current  enters  the 

Cathode  is  the  electrode  by  means  of  which  the  current  leaves  the 

Ions  are  the  constituent  elements  or  radicals  which  carry  the  current 
of  electricity  through  the  electrolyte. 

Anions  are  the  elements  or  radicals  which  appear  at  the  anode. 

Cathions  are  the  elements  or  radicals  which  appear  at  the  cathode. 

Electrolyzer  is  the  apparatus  by  means  of  which  or  in  which  the 
electrolysis  takes  place. 

Diaphragm  is  a  partition,  permeable  or  impermeable,  between  the 
anode  and  cathode,  which  permits  the  passage  of  the  electric  current  but 
prevents  the  mixing  of  the  electrolytes  in  which  the  anodes  and  cathodes 
are  immersed. 

Anolyte  is  the  electrolyte  in  the  anode  compartment  of  the 

Catholyte  is  the  electrolyte  in  the  cathode  compartment  of  the 

Current  Density  is  the  quantity  of  current,  i.e.,  the  number  of  am- 
peres, flowing  through  a  unit  of  electrode  surface. 

Current  Efficiency  is  determined  by  the  yield  per  ampere. 

Energy  Efficiency  is  determined  by  the  decomposition  per  watt. 

Watt. — A  watt  is  the  product  of  one  ampere  multiplied  by  one  volt. 

Horse-power  is  the  equivalent  of  746  watts,  and  is  approximately 
equal  to  3/4  kilowatt. 

Kilowatt  is  1000  watts,  and  is  approximately  equal  to  1  1/3  h.  p. 

Columb. — One  ampere  of  electricity  flowing  for  one  second. 



Electrolysis  has  been  made  the  basis  of  a  number  of  processes  of  ex- 
tracting copper  from  its  ores.  Ever  since  success  was  achieved  in  elec- 
trolytic refining  of  blister  copper,  metallurgists  have  naturally  asked  why 
similar  operations  could  not  be  successfully  applied  in  the  extraction  of 
copper  direct  from  its  ores.  At  first  thought  the  matter  seems  simple 
enough,  but  there  are  difficulties  in  the  way  of  ej^tracting  copper  from  its 
ores  which  are  not  encountered  in  electrolytic  refining.  These  diffi- 
culties, however,  do  not  appear  to  be  insurmountable,  and  it  is  quite 
probable  that  electrolytic  methods  of  extraction  will  be  in  general  use 
in  the  near  future.  Many  of  the  difficulties,  first  met  with,  have  been 
surmounted,  and  the  remaining  ones  are  gradually  being  overcome. 

Anode. — One  of  the  fundamental  differences  between  electrolytic 
refining  and  electrolytic  extraction,  lies  in  the  anode.  In  electrolytic 
refining  the  anode  is  a  slab  of  blister  copper,  usually  about  an  inch  thick, 
and  containing  less  than  1  per  cent,  of  foreign  matter.  This  copper 
anode  goes  into  solution  and  is  redeposited  on  the  cathode.  Theoretic- 
ally, no  energy  is  required  to  perform  this  work,  since  the  energy  de- 
veloped in  dissolving  the  copper  anode  is  equivalent  to  that  consumed 
in  its  deposition  on  the  cathode.  The  soluble  anode  presents  no  serious 
difficulty.  It  has  to  be  replaced  at  frequent  intervals.  Theoretically, 
no  acid  is  consumed  and  none  generated  in  electrolytic  refining,  since  the 
acid  going  into  combination  with  the  copper  at  the  anode  is  again  re- 
leased as  free  acid  by  its  deposition  at  the  cathode. 

When  copper  is  dissolved  from  the  ore,  conditions  are  entirely  differ- 
ent. The  problem  then  is,  to  deposit  the  copper  out  of  solution  by 
electrolysis  while  none  is  going  into  solution.  This  manifestly  involves 
the  use  of  an  anode  which  will  conduct  the  electricity  into  the  copper 
solution,  while  at  the  same  time  the  anode  itself  remains  unattacked,  or 
insoluble.  This  presents  the  first  serious  difficulty  in  precipitating 
copper  from  leaching  solutions. 

It  has  been  found  a  most  difficult  matter  to  provide  a  substance 
which  will  be  a  good  conductor  of  electricity  and  not  be  attacked  by  the 
combined  action  of  the  current  and  the  solvent.  Many  substances  which 
are  sufficiently  permanent,  have  too  high  a  resistance,  and  consequently 
the  power  required  to  overcome  this  resistance  in  the  anode,  makes  the 
process  so  expensive  as  to  be  prohibitive;  while  on  the  other  hand,  sub- 
stances which  have  a  good  electrical  conductivity  are  not  sufficiently 
permanent.  For  chloride  solutions  this  problem  of  suitable  anodes  has 
been  satisfactorily  solved  by  the  use  of  graphitized  carbon,  but  for 
sulphate  solutions  no  really  satisfactory  anode  has  yet  been  discovered, 
although  most  conceivable  substances  have  been  tried.  Lead,  on  the 
whole,  has  given  the  best  results  for  sulphate  solutions.  The  purity  of 
the  lead  is  an  important  factor. 

Platinum  is  too  costly  for  any  process  operated  on  a  working  basis. 


All  other  commercial  metals  are  attacked,  and  even  platinum  is  not 
entirely  unaffected.  Ferro-silicon,  which  is  a  difficultly  attackable 
substance,  has  been  repeatedly  suggested,  but  has  not  proved  successful 
in  practice.  Antimonial  lead  has  been  tried  for  sulphate  solutions,  but 
there  is  no  evidence  to  indicate  that  it  is  more  permanent  than  pure  lead; 
the  antimony  from  the  lead,  going  into  solution  in  the  electrolyte,  may 
be  objectionable  and  its  loss  by  the  decomposition  of  the  anode,  may  be 
a  serious  item  of  expense.  In  almost  all  cases  carbon  is  the  only  sub- 
stance which  can  be  employed  in  chloride  solutions,  but  for  sulphate 
solutions  it  is  absolutely  worthless. 

The  quality  of  carbons  for  electrolytic  work  varies  considerably,  but 
even  the  best  are  eventually  destroyed.  Graphitized  carbons  have  given 
satisfactory  results  with  chloride,  but  not  with  sulphate,  solutions. 
These  carbons,  in  addition  to  being  durable,  are  good  conductors  of 
electricity.  They  possess  a  specific  resistance  of  but  0.00032  ohm  per 
cubic  inch,  which  is  only  one-fourth  that  of  amorphous  carbon.  The 
specific  resistance  is  0.000813  ohm  per  cubic  centimeter,  and  since  that 
of  mercury  is  0.000094  ohm,  the  conductivity  is  11.75,  mercury  being 
100;  or  0.21,  copper  being  100.  In  some  cases  graphitized  carbons  have 
been  used  for  3  years  as  anodes  in  the  decomposition  of  alkali  metal 
chloride  solutions,  with  a  current  density  of  50  to  250  amperes  per  square 
meter.  It  is  essential,  however,  that  the  solution  be  acid;  in  alkaline  or 
neutral  solutions,  they  are  not  as  durable.  The  following  is  a  list  of 
standard  sizes  and  weights  of  Acheson-Graphite  electrodes: 

Acheson-Graphite  Electrodes 









n.  diam.  X12  in., 
in.  diam.  X 12  in., 
n.  diam.  X24  in., 
in.  diam.  X  24  in., 
n.  diam.  X 24  in., 
n.  diam.X24  in., 
in.  diam.  X  24  in., 
in.  diam.  X  24  in., 
in.  diam.  X 24  in., 
n.  diam.  X 24  in., 
n.  diam.  X  24  in., 
n.  diam.  X 24  in., 
n.  diam.  X  24  in., 
n.  diam.  X  40  in., 
n.  diam.  X 40  in., 
in.  diam.  Xl9i  in., 
n.  diam.  X 48  in., 
in.  diam.  X 48  in.. 

Approximate  weight 
per  piece 

0.008  lb. 

0.035  1b. 

0.15  lb. 

0.21  lb. 

0.26  lb. 

0.40  lb. 

0.60  lb. 

0,82  lb. 

1,00  lb. 

1,40  lb. 


2,60  lb. 

4,50  lb. 
16.60  1b. 
30.00  lb. 
23.50  lb. 
75.00  lb. 
132.00  !b. 

Approximate  wei 

per  piece 


.50  1b. 


,00  lb. 


,50  1b. 


,  10  lb. 

2.94  1b. 

4.50  1b. 

4.05  lb. 


.13  lb. 


.21  lb. 


.75  1b. 


.20  1b. 


.  50  lb. 


.30  1b. 


.88  lb. 


.50  1b. 


50  lb. 


.75  lb. 


,25  1b. 


.45  lb. 


Acheson-Graphite  Electrodes. — Continued 


2X2X30  in., 
4X4X40  in., 
6X6X40  in.. 

J  X 12X12  in., 

iX4   X24in., 

iX6   X24in., 

iX  12X12  in., 

}X Six  19 J  in., 

fX5   XlSin., 

1x12X12  in., 

|X2   X21-Hn., 
1    X4   X30in., 

1  X6  X30in., 
liX3  X36in., 
liX5   X30in., 

2  X4  X30  in., 

2  X7   XSOin., 

3  X6   XSOin., 

4  X8fXl5in., 

These  electrodes  cost  approximately  from  $13.50  to  $15.00  per  100  lb. 
in  lots  of  500  lb.  for  the  sizes  most  convenient  for  electrolytic  use.  In 
ton  lots  the  cost  would  be  approximately  $11.50  per  100  lb. 

Cathodes. — The  cathodes  used  in  electro  deposition  are  usually 
thin  sheets  of  pure  copper.  Carbon  cathodes,  in  chloride  solutions,  do 
very  well  but  are  not  advisable  for  sulphate  solutions.  If  copper  cathode 
sheets  are  used,  they  may  be  stripped  after  the  deposit  has  acquired  the 
desired  thickness,  or  new  sheets  may  be  supplied  as  the  old  ones  are 
removed.  Lead  cathode  sheets  have  been  used,  and  the  copper  stripped 
from  the  lead,  or,  the  lead  may  be  melted  from  the  copper  after  removal 
from  the  electrolyzers,  and  again  rolled  into  sheet  lead  for  new  cathode 

In  depositing  copper  from  impure  solutions,  that  is  to  say,  solutions 
like  those  obtained  in  leaching  copper  ores,  it  is  not  an  easy  matter  to 
get  a  reg-uline  deposit  of  the  desired  thickness.  Unless  considerable  care 
is  taken  with  the  electrolyte  and  the  current  density,  irregular  deposition 
and  sprouting  will  occur  long  before  the  cathodes  have  acquired  the 
thickness  desired  for  their  removal.  This  difficulty  may  be  so  aggravated 
as  to  make  their  removal  necessary  at  an  early  stage  of  the  operation,  and 
thus  adding  considerable  to  the  expense.  If  the  cathodes,  under  such 
conditions,  are  not  removed,  the  difficulty  is  quickly  aggravated  and 
short  circuiting  and  inefficiency  are  likely  to  result.  With  a  reasonably 
pure  solution  and  low  current  density,  this  difficulty  is  not  likely  to  occur. 


especially  if  the  solution  is  agitated  or  if  the  cathode  is  moved  through  the 
electrolyte.  The  current  density,  as  well  as  the  nature  of  the  electrolyte, 
has  much  to  do  with  the  quality  of  the  deposit;  the  lower  the  current 
density,  the  more  reguline  the  deposit  is  likely  to  be,  but  there  is  a  mini- 
mum practical  limit  to  the  current  density  that  can  be  employed. 

Diaphragms.— Most  of  the  electrolytic  processes  are  based  on  the  fact 
that  the  solvent  may  be  regenerated  during  electrolysis.  If  the  solution 
electrolyzed  at  the  cathode  requires  to  be  kept  separated  from  that  at 
the  anode,  then  diaphragms  are  necessarily  inserted  between  the  anode 
and  cathode,  which,  while  allowing  the  free  passage  of  the  current,  will 
prevent  the  solutions  from  mixing.  Or  if  the  electrolyte  contains  a 
large  amount  of  an  oxidizable  and  reducible  metal  under  the  anodic  and 
cathodic  influences,  it  may  be  desirable  to  use  diaphragms,  simply  to 
overcome  the  undue  loss  in  electrical  efficiency.  If  copper  sulphate 
solution,  containing  much  iron  sulphate  is  electrolyzed,  considerable 
energy  may  be  consumed  in  the  oxidation  and  reduction  of  the  iron, 
which  simply  renders  an  equivalent  in  useless  heat. 

Various  materials  may  be  used  for  diaphragms,  but  not  many  fulfill 
the  conditions  of  low  electrical  resistance  combined  with  sufficient  density 
to  prevent  the  anode  and  cathode  solutions  from  mixing  too  freely.  The 
material  best  suited  for  diaphragms  is  asbestos,  which  is  not  easily  attacked 
either  by  acid  or  alkaline  solutions,  and  when  saturated  with  the  electro- 
lyte, offers  no  appreciable  resistance  at  low  current  densities.  At  high 
current  densities,  any  diaphragm  is  likely  to  offer  appreciable  resistance. 
Asbestos,  suitable  for  diaphragms,  is  manufactured  either  as  cloth,  paper, 
or  mill  board.  Asbestos  cloth  will  give  the  best  results  if  it  is  not  neces- 
sary to  make  an  absolute  separation  between  the  anolyte  and  catholyte. 
When  complete  separation  is  necessary  it  is  desirable  to  use  mill  board  or 
asbestos  paper  in  connection  with  asbestos  cloth.  In  whatever  form  the 
asbestos  is  used,  it  is  desirable  that  it  be  quite  free  from  foreign  matter 
or  admixtures. 

While  the  construction  of  suitable  diaphragms  is  not  an  insurmount- 
able difficulty,  it  is  desirable  to  dispense  with  them  wherever  practicable. 
The  insertion  of  diaphragms  in  the  electrolyzer  also  presents  difficulties, 
as  well  as  the  construction  of  the  diaphragm  itself.  In  a.  copper  electrolyte 
the  ordinary  materials  of  construction  cannot  be  used,  and  hence  to 
employ  materials  to  take  the  place  of  the  ordinary  iron  nails  and  bolts  is 
more  or  less  expensive.  Ordinary  porous  clay,  such  as  is  used  in  battery 
jars  makes  a  good  diaphram  to  keep  the  solutions  separate,  but  its  elec- 
trical resistance  is  too  high  to  be  used  in  practice,  and  it  must  be  used  in 
sizes  too  small  for  economic  adaptation. 

It  is  evident  that  a  diaphragm  if  it  is  to  fulfill  its  purpose,  its  solid 
particles  must  not  carry  any  current,  since  otherwise  the  diaphragm 
would  not  act  as  a  diaphragm,  but  as  a  bipolar  electrode.     The  current 


passes  through  the  interstices  of  the  solid  particles  which  compose  the 
diaphragm,  and  the  resistance  of  the  diaphragm  is  the  composite  resistance 
of  all  the  innumerable  passageways  of  the  electrolyte  through  the  pores 
of  the  diaphragm.  Diaphragms  should  be  easily  permeable  to  charged 
ions  but  should  offer  a  high  resistance  to  diffusion. 

In  Denver,  at  the  Greenawalt  experimental  plant,  were  used  some 
years  ago,  on  a  large  scale,  diaphragms  made  of  asbestos  cloth,  asbestos 
paper,  or  mill  board,  or  a  combination  of  these,  sandwiched  between 
perforated  boards  from  1/2  to  3/4  in.  thick.  The  perforations  were  about 
1/2  in.  in  diameter  and  as  close  as  possible,  consistent  with  strength.  The 
boards,  with  the  asbestos  between,  were  fastened  together  with  wooden 
dowel  pins  and  keyed.  These  diaphragms  were  expensive,  and  as  the 
perforations  did  not  exceed  half  the  total  area,  only  about  half  the 
diaphragm  was  effective.  Diaphragms  constructed  in  this  way,  must  of 
necessity  be  small.     Their  use  was  discontinued. 

Later,  W.  E.  Greenawalt  constructed  diaphrams  which  were  used  in 
copper  electro  deposition,  and  were  built  13  ft.  6  in.  long  by  3  ft.  wide, 
made  of  two  sheets  of  asbestos  cloth,  between  which  were  sandwiched 
the  desired  thickness  of  asbestos  paper,  and  these  in  turn,  were  sand- 
wiched between  two  mullioned  oak  frames,  bolted  together  with  copper 
bolts.  These  diaphragms  gave  satisfactory  results,  but  were  later  dis- 
carded when  it  was  found  that  for  the  process  under  demonstration,  dia- 
phramgs  could  be  dispensed  with  entirely. 

Betts  states  that  the  best  diaphragm  he  knows  of  for  ordinary  weak 
acid  solutions  may  be  made  as  follows:'  Powdered  sulphur  is  sifted 
evenly  over  a  1/4-in.  asbestos  mill  board  and  the  sheet  heated  evenly 
an  hour  or  so  just  above  the  melting  point  of  sulphur.  The  operation 
is  then  repeated  with  the  other  side.  It  takes  from  1/4  to  3/4  lb.  of 
sulphur  to  the  square  foot  of  mill  board.  This  diaphragm  is  plastic 
when  heated  and  in  solutions  does  not  soften  at  all.  It  expands 
slightly,  and  should  be  kept  in  acid  water  two  or  three  weeks  before 
using  in  any  kind  of  rigid  construction.  The  resistance  is  somewhat 
higher  than  without  the  sulphur,  but  i,he  diffusion  is  smaller. 

Current  Density. — The  current  density  has  much  to  do  with  the  effi- 
ciency of  the  operation  and  the  nature  of  the  copper  deposited.  It  is 
not  possible,  in  electrolyzing  impure  leaching  solutions,  to  use  as  high  a 
current  density  as  in  electrolytic  refining,  owing  principally  to  the  im- 
purities, and  to  some  extent  the  leaner  copper  content.  If  a  reguline 
deposit  is  desired,  it  will  be  necessary  to  a  rather  low  current  density. 
In  electrolytic  decomposition  high  current  density  causes  impoverish- 
ment of  ions  at  the  electrodes  and  causes  trouble.  This  impoverishment 
is  prevented  by  stirring  the  electrolyte  or  moving  the  electrodes. 

The  energy  efficiency  becomes  less  as  the  current  density  is  increased, 

'  "  Electrochemical  Industry,"  July,  1908. 


and  the  theoretical  voltage  is  only  approached  at  very  low  current 
densities.  If  the  voltage  for  a  certain  electrolyte,  for  example,  is  1.2  at 
5  ampcics  per  square  foot,  at  a  current  density  of  30  amperes  per 
square  foot  it  is  very  likely  to  be  three  times  that.  Or,  in  other  words, 
the  energy  efficiency  at  a  current  density  of  30  amperes  per  square  foot, 
is  likely  to  be  only  one-third  that  at  5  amperes  per  square  foot,  assuming, 
of  course,  that  the  amount  of  copper  deposited  is  the  same  in  both  cases, 
which  may  be  a  gratuitous  assumption.  It  follows,  further,  that  while 
the  electrolytic  deposition  of  the  copper  will  take  three  times  the  power 
at  30  amperes  per  square  foot  as  it  would  for  5  amperes  per  square  foot, 
the  size  of  the  plant  would  be  only  one-sixth  as  large,  and  presumably 
cost  only  one-sixth  as  much. 

The  lower  current  densities  will  be  largely  limited  to  the  cost  and  size 
of  installation  for  the  same  output,  and  the  larger  current  densities  by  the 
cost  of  operation. 

The  cost  of  power  will  be  a  governing  factor  in  deciding  the  current 
density  to  be  employed.  If  power  is  cheap  it  will  probably  cost  no  more 
to  operate  at  a  fair  current  density  than  at  a  low  current  density;  for 
while  the  power,  per  pound  of  copper,  will  cost  more,  this  will  be  offset 
by  the   decreased   cost  of  installation,  maintenance,  and   attendance. 

If  high  current  densities  are  employed,  it  will  be  desirable  to  agitate 
the  electrolyte,  or  employ  some  form  of  moving  cathode,  so  as  to  bring 
sufficient  copper  ions  in  contact  with  the  cathode  and  thus  avoid  useless 
expenditure  of  energy  in  the  decomposition  of  other  substances  in  the 
electrolyte,  or  even  the  electrolyte  itself. 

It  was  observed  soon  after  practical  attempts  were  made  at  copper 
refining,  about  the  year  1865,  that  the  current  density,  and  consequently 
the  rate  of  deposition  could  be  considerably  increased  by  circulating  the 
electrolyte  or  moving  the  electrodes. 

Wilde  was  one  of  the  first  to  deposit  copper  on  a  revolving  cathode. 
The  anodes  consisted  of  copper  cylindrical  tubes,  and  the  cathode  con- 
sisted of  an  iron  cylinder  which  was  to  be  coated  with  copper.  The 
cathode  was  placed  in  the  center  of  the  electrolyzer  and  rotated  on  its 
axis.  This  gave  an  even  distribution  of  the  copper  over  the  entire 
cathode  surface  by  means  of  the  motion  imparted  to  the  solution  and 
the  equal  current  density  resulting  from  the  motion.  The  current 
density  used  was  about  20  amperes  per  square  foot. 

Elmore  used  horizontal  mandrels  on  which  copper  sheets  or  tubes 
are  deposited,  while  agate  burnishers  travel  continuously  over  the  copper 
so  as  to  consolidate  it  and  at  the  same  time  prevent  the  growth  of  copper 
trees  or  nodules.  The  current  density  used  was  about  30  amperes  per 
square  foot. 

Dumoulin  introduced  a  process  for  burnishing  copper,  during  deposi- 
tion, with  sheepskin  as  a  substitute  for  agate.     He  claimed  that  the 



process  had  the  additional  advantage  of  insulating  any  projections  that 
might  be  formed  on  the  deposited  metal.  It  was  claimed  that  a  current 
of  from  30  to  40  amperes  per  square  foot  of  cathode  surface  could  be  used 
with  a  difference  of  potential  of  1.6  volts,  using  a  soluble  copper  anode. 

Cowper-Coles  gets  excellent  deposits  of  any  desired  thickness  by 
revolving  a  cylinder  cathode  at  a  speed  of  from  1500  to  2000  lin.  ft.  per 
minute,  with  a  current  density  of  200  amperes  per  square  foot. 

Attempts  have  been  made  at  various  times  to  increase  the  rate  of  cop- 
per deposition  by  Swan,  Elmore,  Thofern,  Graham,  Poore,  and  others  by 
impinging  jets  of  the  electrolyte  against  a  cathode  surface.  The  quality 
of  the  deposits  is  likely  to  be  unsatisfactory  if  impinging  jets  are  alone 
employed;  it  is  desirable,  therefore,  to  move  the  cathode  also,  in  order 
that  the  deposited  copper  may  be  uniform  over  its  entire  surface. 

Whatever  the  means  employed  to  increase  the  rate  of  deposition,  the 
essential  object  to  be  attained  is  to  bring  sufficient  copper  ions  in  contact 
with  the  cathode,  corresponding  to  the  increased  current,  and  to  get  a 
reguline  deposit  by  friction  either  with  the  electrolyte  or  burnisher. 

General  Laws  Governing  the  Electrodeposition  of  Copper. — In  the 
decomposition  of  copper  solutions,  work  is  performed  and  energy  con- 
sumed. The  factors  governing  the  expenditure  of  energy,  in  electrolysis, 

The  electromotive  force. 
The  resistance,  and. 
The  current. 

The  electromotive  force  (e.  m.  f.)  is  measured  in  volts;  the  resistance 
in  ohms,  and  the  current  in  amperes.  A  definite  relation  exists  between 
these  three  factors,  whereby  the  value  of  any  factor  may  always  be  calcu- 
lated when  the  value  of  the  other  two  are  known.  This  relation  is  known 
as  Ohm's  law,  and  may  be  stated  thus: 

The  current  strength  in  any  circuit  is  equal  to  the  electromotive  force 
applied  to  the  circuit,  divided  by  the  resistance  of  the  current.  Or  more 
briefly  stated: 


Current     = 


Amperes  =   -—- —  or 

C        = 



In  other  words,  the  electromotive  force  (e.  m.  f.)  which  may  be 
assumed  to  be  the  electrical  pressure  by  the  dynamo,  causes  the  flow  of  an 
electric  current.  The  current  is  directly  proportional  to  the  electromo- 
tive force.     The  resistance  (electrical  conductors  and  electrolyte)  oppose 


tlic  flow  of  the  curjcnl.     The  current  is  therefore  proportional-  to  the 
electromotive  force. 

The  current  strength  in  any  circuit  increases  or  decreases  directly 
as  the  electromotive  force  increases  or  decreases,  when  the  current  is 
constant.  With  a  constant  pressure  the  current  increases  as  the  resist- 
ance is  decreased,  and  decreases  as  the  resistance  is  increased;  or  briefly, 
the  current  varies  directly  as  the  electromotive  force  and  inversely  as  the 

From  this  it  follows,  that  to  double  the  resistance  halves  the  current, 
the  electromotive  force  remaining  constant.  Or,  if  with  the  doubled 
resistance  the  current  is  to  remain  constant,  the  electromotive  force, 
and  consequently  the  power,  must  be  doubled. 

It  also  follows  from  Ohm's  law  that  when  through  a  given  resistance 
the  current  is  required  to  be  doubled,  the  power  must  be  increased  four 
times;  or  in  general,  the  resistance  remaining  constant  the  power  in- 
creases proportionately  to  the  square  of  the  current. 

Ohm's  law,  applied  to  the  deposition  of  copper,  means  that,  theoretic- 
ally, the  least  power  is  required  to  deposit  a  given  amount  of  copper 
when  the  lowest  possible  current  density  is  used.  This  would  require  in 
practice,  a  plant  of  unlimited  size,  so  that  a  well  designed  plant  must  of 
necessity  be  a  compromise  between  theoretical  conditions  and  practical 

The  Electrical  Power  is  the  product  of  the  number  of  volts  multiplied 
by  the  number  of  amperes  of  the  current.  Its  unit  is  the  Watt.  746 
watts  make  a  horse  power;  1000  watts  a  kilowatt.  Either  the  horse 
power  or  the  kilowatt  may  be  taken  as  the  unit  of  power;  the  kilowatt 
is  the  most  convenient  in  electrical  and  electrolytic  work. 

The  power  consumed  in  depositing  a  definite  amount  of  copper  may 
be  expressed  in  horse  power,  thus: 

Volts  X  Amperes 

^-  P-  = 746 

and  in  kilowatts, 

Volts  X  Amperes 

k.  w.  = 


The  most  important  laws  in  relation  to  electrolyais  are  those  of 
Faraday.     His  law  of  electrolysis  may  be  stated  thus: 

Faraday's  Law. — 1.  The  amount  of  chemical  change  produced  electro- 
lytically  by  the  current  is  proportional  only  to  the  amount  of  electricity 
passing,  as  measured  in  columbs,  and  is  independent  of  the  strength  or 
temperature  of  the  electrolyte,  or  the  size  or  distance  apart  of  the  electrodes. 

2.  The  amount  of  different  elements  dissolved  or  set  free  by  the  passage 
of  a  given  amount  of  electricity  is  proportional  to  their  chemical  equivalents. 
Or  in  other  words,  the  amount  of  electrochemical  action   produced  is 


directly  proportional  to  the  product  of  ampere  hours  and  the  chemical 

A  given  quantity  of  current,  therefore,  will  always  deposit  the  same 
quantity  of  a  given  element,  and  the  elements  are  deposited  propor- 
tionally to  their  chemical  equivalents;  as  for  example,  hydrogen  1, 
oxygen  8,  chlorine  35.5,  cupric  copper  31.75,  cuprous  copper  63.5,  etc. 

_,,      ,,  ^,       .    ,  ^     .     ,       ,,  .      ,  ,.        Atomic  weight     ^, 

ihe      Chemical  Equivalent      is  the  quotient  -, ihe 

^  ^  valency 

electrochemical  equivalent  of  a  substance  is  identical  in  weight  with  the 

amount  of  the  same  substance  that  would  be  deposited  by  1  ampere 

in  1  second  (that  is,  1  columb).     The  electrochemical  equivalent  of  any 

element  may  be  calculated  by  multiplying  the  chemical  equivalent  of 

the  element  by  .000010384,  which  is  the  electrochemical  equivalent  of 

hydrogen.     For  cupric  copper  it  is  0.0003297  and  for  cuprous  copper 

it  is  0.0006594  (grams  per  columb). 

Theoretical  Data  for  Copper  Deposition. — Theoretically  1.1858  grm. 
of  copper  are  deposited  by  1  ampere-hour  from  cupric,  and  2.3717  grm. 
per  ampere-hour  from  cuprous  solutions.  This  for  practical  purposes 
would  mean  that  1  ampere  should  deposit  1  oz.  of  copper  from  cupric 
solutions,  and  2  oz.  from  cuprous  solutions,  per  day  of  24  hours.  From 
this  it  is  easy  to  compare  the  current  efficiency  of  any  process  or  opera- 
tion with  the  theoretical  amount,  which  may  be  taken  as  100  per  cent. 

The  following  theoretical  depositions  of  copper,  referred  to  various 
units,  will  Ijc  found  convenient  for  reference: 

One  ampere-hour  will  deposit,  theoretically,  1.1858  grm.  of  copper 
from  cupric  solutions,  and  2.3717  grm.  from  cuprous  solutions. 

0.8433  ampere-hours  will  deposit  1  grm.  of  copper  from  cupric  solu- 
tions, while  0.42164  ampere-hours  will  deposit  1  grm.  from  cuprous 

One  ampere-hour  will  deposit  0.02614  lb.  of  copper  from  cupric  solu- 
tions and  0.05228  lb.  from  cuprous  solutions. 

382.50  ampere-hours  will  deposit  1  lb.  of  copper  from  cupric  solutions; 
191.25  ampere-hours  will  deposit  a  pound  of  copper  from  cuprous 

746  ampere-hours  will  deposit  1.9494.1b.  of  copper  from  cupric,  and 
3.8988  lb.  from  cuprous  solutions. 

1000  ampere-hours  will  deposit  2.6143  lb.  of  copper  from  cupric,  and 
5.229  from  cuprous  solutions. 

These  statements  of  deposition  do  not  take  into  consideration  the 
voltage  at  which  the  current  is  delivered.  It  simply  represents  the 
amount  of  copper  irrespective  of  the  voltage.  While  the  current  efficiency 
of  a  process  may  be  determined  from  the  above  theoretical  quantities, 
the  energy  efficiency  can  only  be  determined  when  the  theoretical  as  well 
as  the  practical  voltage  is  known. 



The  theoretical  voltage  is  proportional  to  the  heats  of  combination  of 
the  compounds  decomposed,  and  may  be  calculated  from  the  molecular 
heats  of  combination  as  a  basis.  For  cupric  sulphate  it  is  1.22  volts;  for 
cupric  chloride  it  is  1.35  volts,  and  for  cuprous  chloride  it  is  1.42  volts. 
Blount'  found  by  actual  experiment  in  his  laboratory  that  the  minimum 
■pressure  necessary  for  the  deposition  of  copper  from  cupric  sulphate, 
using  an  insoluble  anode,  is  1.375  volts. 

Knowing  the  theoretical  deposition  of  copper  by  a  definite  amount 
of  electric  current,  and  the  theoretical  voltage  as  determined  from  the 
heats  of  combination,  the  theoretical  power  consumed  in  electrodeposi- 
tion  may  readily  be  calculated,  and  this  is  taken  as  the  standard  of  the 
energy  efficiency,  or  100  per  cent.,  to  which  all  practical  energy  efficiencies 
may  be  referred,  since  that  represents  the  greatest  possible  amount  of 
copper  that  can  be  deposited  with  a  given  amount  of  electrical  energy. 
It  represents  the  greatest  amount  of  copper  that  can  be  deposited  with  a 
definite  current,  and  the  lowest  possible  voltage  at  which  it  can  be 

The  following  tabulated  statement,  gives  in  convenient  form,  the 
theoretical  energy  required  to  deposit  copper,  with  insoluble  anodes, 
from  cupric  sulphate,  cupric  chloride,  and  cuprous  chloride  solutions, 
based  on  the  molecular  heats  of  combination. 


Cupric  sulphate. . . 
Cupric  chloride. . . 
Cuprous  chloride.. 

Grm.  per 
h.  p. -hour 




Grm.  per 

Lb.  per 

Lb.  per 

Lb.  per 

c.  w.-hour 

h.  p. -hour 

k.  w.-hour 

h.  p.  day    1 





1 .  9440 


Lb.  per 
■h.  p.  day 


88 .  38 

To  determine  the  actual  current  and  energy  efficiencies  of  copper  depo- 
sition in  an  electrolytic  process,  the  current  is  accurately  measured  both 
as  to  quantity  and  pressure — amperes  and  volts — for  a  certain  definite 
time.  The  amperes  multiplied  by  the  volts,  gives  the  watts;  746  watts  is 
equal  to  a  horse-power,  and  1000  watts  to  a  kilowatt.  The  copper  de- 
posited is  carefully  collected,  dried,  weighed  and  assayed  to  determine 
its  purity.  The  weight  of  the  pure  copper  deposited  by  the  current  and 
the  power  consumed,  compared  with  the  theoretical,  gives  the  current 
and  energy  efficiencies. 

For  Example. — In  making  a  certain  test  on  the  electrodeposition  of 
copper  from  a  cuprous  solution,  a  current  of  400  amperes  was  used  foi 
12  hours  (4800  ampere-hours),  at  1.8  volts.  The  pure  copper  recovered 
was  18.2  lb.  The  theoretical  amount  that  should  have  been  deposited 
at  the  rate  of  5.229  lb.  per  1000  ampere-hours,  is  25.1  lb.     The  test,  there- 

'  "Practical  Electro-Chemistry,"  p.  64,  1906. 


fore,  showed  a  current  efficiency  of  72.5  per  cent.  The  copper  deposited 
in  the  test  was  at  the  rate  of  2.15  lb.  per  k.  w.-hour;  the  theoretical 
deposition  per  k.  w.-hour  is  3.6824  lb.;  hence  the  energy  efficiency  is 
58.2  per  cent.  As  compared  with  the  deposition  from  cupric  solutions, 
however,  this  test  would  show  a  current  efficiency  of  145  per  cent,  and 
an  energy  efficiency  of  116.4  per  cent. 

In  practical  electrolytic  work  it  is  rarely  that  90  per  cent,  is  exceeded 
for  the  current  efficiency,  and  50  per  cent,  for  the  energy  efficiency.  It 
will  usually  be  found  profitable  to  sacrifice  electrical  efficiency  for  other 

In  many  processes,  the  secondary  anode  reactions  tend  to  reduce  the 
theoretical  voltage;  this  may  vitally  effect  the  energy  efficiencies  as  above 
given,  because  the  secondary  reactions  must  be  taken  into  consideration 
in  figuring  the  eneTgy  efficiency  for  any  particular  process,  when  such 
reactions  are  involved.  In  the  Hoepfner  process,  for  example,  the 
theoretical  voltage  for  the  decomposition  of  cuprous  chloride  is  1.42 
volts,  but  this  is  to  some  extent  offset  by  the  recombination  of  the  released 
chlorine  at  the  anode,  combining  with  the  cuprous  chloride  to  form 
cupric  chloride,  so  that  the  theoretical  voltage  is  only  0.18  volt,  instead 
of  1.42  volts,  or  only  about  one-eighth.  The  theoretical  deposition,  tak- 
ing into  account  the  secondary  reaction,  would  be  697.2  lb.  of  copper  per 
k.  w.-day  of  24  hours  instead  of  88.38  as  given  in  the  table.  A  similar 
reduction  of  voltage  is  theoretically  possible  in  the  Siemens-Halse  process 
as  also  in  processes  using  sulphur  dioxide  to  combine  with  chlorine  or 
sulphion  at  the  anode.  It  must  be  evident,  however,  that  these  reduc- 
tions of  voltage  due  to  secondary  reactions,  are  more  apparent  than  real, 
largely  because  an  impracticable  small  current  density  must  be  employed 
for  their  realization,  or  even  an  approach  at  realization,  although  Hoepf- 
ner claimed  a  practical  voltage  of  0.8  volt  in  the  operation  of  his  process; 
but  Hoepfner  used  an  exceedingly  low  current  density. 

Loss  of  Energy  in  Electrolytic  Work:  Joule's  Law. — Much  of  the 
energy  consumed  in  electrolytic  work,  both  in  the  conductors  and  electro- 
lyte, appears  as  heat.  Joule  first  discovered  that  the  development  of 
heat  was  proportional  to: 

1.  The  resistance  of  the  conductor; 

2.  The  square  of  the  current; 

3.  The  time  during  which  the  current  flows. 

When  heat  is  thus  produced  in  the  conductors  and  the  electrolyte  it 
is  a  waste  of  power.  The  current  density  both  in  the  electrolyte  and  in 
the  conductors  may  seriously  affect  the  heating  of  the  circuit,  and  con- 
sequently the  ultimate  voltage  at  the  dynamo.  The  conductors  should, 
therefore,  be  amply  large  to  carry  the  current,  and  the  electrical  connec- 
tions should  be  as  few  as  possible,  and  all  connections  should  be  well 


made.  At  the  Anaconda  Copper  Refinery  it  was  found  on  careful  in- 
vestigation, that  20  per  cent,  of  the  loss  of  efficiency  was  in  the  conductor 

A  considerable  percentage  of  the  power  used  in  electrodeposition  of 
the  copper  may  be  lost  in  the  metallic  conductors,  and  contacts,  especially 
if  the  conductors  are  of  insufficient  cross  section  and  the  connections  poor. 

For  electrolytic  refineries,  which  are  presumably  under  at  least  good 
average  conditions,  Addicks'  gives  a  rough  summary  of  the  relative  value 
of  the  resistances  in  practice,  as  follows: 

Metallic  resistance,  15  per  cent. 

Electrolyte,  including  transfer,  60  per  cent. 

Contacts,  20  per  cent. 

Counter  e.  m.  f.,  5  per  cent. 

Contact  resistances  are  met  with  at  the  joints  in  the  main  bars  and  at 
the  connections  between  the  bars  and  the  electrodes.  The  joints  in  the 
main  bars  should  be  equal  in  conductivity  to  the  bar  itself.  This  stand- 
ard can  be  easily  attained  if  the  bars  are  properly  faced.  Three  hundred 
to  four  hundred  amperes  per  square  inch  of  bearing  area  will  give  no 

The  counter  electromotive  force  in  copper  refining,  due  to  the  greater 
concentration  of  the  electrolyte  at  the  anode  than  at  the  cathode  is  quite 
small,  usually  about  0.02  volt.  Contacts  for  copper  deposition  from 
insoluble  anodes  are  more  likely  to  be  source  of  loss  of  energy  than  in 
electrolytic  refining,  due  to  the  considerably  higher  voltage  between  the 
electrodes.  In  copper  refining  the  difference  of  potential  between  the 
electrodes  is  from  0.2  to  0.4  volt,  while  in  copper  deposition  with  insolu- 
ble anodes,  it  will  vary  in  practice  between  1.5  and  3  volts,  depending 
largely  on  the  current  density  employed. 

The  Electrolyte. — The  solvents  usually  employed  for  electrolytic 
processes  have  either  sulphuric  or  hydrochloric  acid  as  the  basis.  The 
solvent,  to  a  large  extent,  determines  the  details  of  the  process.  In  the 
electro  deposition,  the  solvent  should  be  regenerated,  as  well  as  the  cop- 
per precipitated.  In  leaching  copper  ores,  therefore,  repeatedly  with 
the  same  solution  all  the  soluble  impurities  in  the  ore  are  likely  to  be  found 
in  the  electrolyte,  and  this  is  true,  but  not  to  the  same  extent,  if  the  solu- 
tion is  used  only  once  and  then  wasted,  instead  of  being  used  repeatedly 
in  the  same  cycle.     The  cycle  will  ordinarily  consist  of: 




'  The  Journal  of  the  Franklin  Institute,  Dec,  1905. 


and  it  is  not  liliely  that  any  electrolytic  process  can  achieve  marked 
success  on  any  other  basis.  The  regeneration  may  be  direct;  that  is  to 
say,  take  place  in  the  electrolyzer  and  under  the  action  of  the  current ;  or, 
the  anode  products  may  be  withdrawn  from  the  electrolyzer  and  then  in 
some  way  combined  with  the  solution. 

Whatever  method  is  adopted,  the  solution,  in  a  cyclic  process,  is  likely 
to  become  charged  with  impurities  and  thus  reduce  the  efficiency  of  the 
deposition.  Two  alternatives  present  themselves;  either  to  purify  the 
solution,  or  waste  it  at  intervals.  If  the  solution  is  wasted,  a  small 
portion  of  it  may  be  withdrawn  at  every  cycle,  and  a  corresponding  a- 
mount  of  fresh  water  added.  This  can  be  done  quite  readily  if  there  is 
more  solvent  regenerated  in  the  process  than  is  consumed  by  the  ore. 

\'arious  methods  have  been  suggested  for  purifying  a  foul  electrolyte; 
these  differ  somewhat  as  a  sulphate  or  a  chloride  solution  is  used. 

The  injurious  effect  of  an  impure  electrolyte  is  shown  more  particularly 
in  the  electrolysis.  The  undesirable  metals  in  the  solution  may  be  de- 
posited with  the  copper  if  the  solution  has  become  impoverished,  or  else 
useless  energy  may  be  expended  in  reduction  and  oxidation,  or  in  depo- 
sition and  immediate  solution,  under  the  influence  of  the  current. 

A  comparative  measure  of  the  energy  required  to  deposit  the  various 
metals,  is  obtained  by  taking  the  heat  of  combination  of  the  metals  with 
oxygen,  to  form  salts.  Those  metals  are  first  deposited  which  have  the 
lowest  heat  of  combination,  other  things  being  equal.  The  order  of 
deposition  may  be  approximately  stated  as  follows:  gold,  silver,  copper, 
antimony,  bismuth,  arsenic,  lead,  nickel,  cobalt,  cadmium,  tin,  iron,  zinc, 
manganese.  The  alkali  metal  salts,  either  the  chlorides  or  sulphates,  are 
only  decomposed  with  considerable  difficulty,  nevertheless,  when  high 
current  densities  are  employed  they  may  frequently  be  the  cause  of  con- 
siderable loss  of  efficiency. 

Gold  and  silver  will  be  deposited  with  the  copper,  and  the  deposition 
may  take  place  either  chemically  or  electrolytically.  Neither  gold  nor 
silver  are  likely  to  be  in  sulphate  solutions;  both  may  be  in  chloride 

The  order  of  precipitation,  as  given,  is  dependent  upon  certain  con- 
ditions which  must  be  observed.  Principally  among  these  conditions 
are  the  strength  of  current,  the  nature  of  the  electrolyte,  the  relative 
proportion  of  the  metals  in  solution,  and  if  the  solution  is  low  in  copper, 
whether  or  not  the  electrolyte  is  agitated.  If  the  current  density,  and 
consequently  the  voltage,  exceeds  a  certain  strength,  all  the  metals,  or 
several  of  them,  may  be  deposited  together.  The  more  neutral  the  elec- 
trolyte is,  the  more  easily  will  the  more  electropositive  metals  be  depos- 
ited. The  current  is  always  striving  to  decompose  the  electrolyte  into 
metal  or  oxide  and  acid  or  the  acid  radical;  while  the  liberated  acid  is 
striving  to  redissolve  the  liberated  metal  or  oxide.     These  two  forces  are 


always  opposed  to  one  another,  and  under  varying  conditions  either  may 
gain  the  upper  hand.  The  resolvent  action  of  the  acid,  in  cases  where  the 
components  of  the  electrolyte  have  a  strong  chemical  affinity,  may 
overpower  the  action  of  a  weak  current. 

In  the  deposition  of  copper  this  secondary  reaction  may  not  be  of  much 
importance  in  a  sulphate  electrolyte,  but  it  is  quite  noticeable  in  the  pres- 
ence of  good  circulation  of  the  electrolyte,  and  more  or  less  access  of  air 
to  the  cathodes.  With  a  chloride  electrolyte  this  secondary  reaction  may 
be  quite  pronounced  under  certain  conditions,  especially  if  the  electrolyte 
is  violently  agitated.  With  all  of  the  easily  dissolvable  metals,  it  is  this 
secondary  reaction  which  prevents  the  metal  from  making  its  appearance 
at  the  cathode  in  metallic  form,  as  the  deposition,  and  solution  due  to 
secondary  reaction,  are  practically  simultaneous,  and  the  operation 
simply  results  in  a  corresponding  loss  of  efficiency. 

It  is  evident  that  if  the  current  density  exceeds  in  amount  that  capable 
of  being  supplied  with  copper  ions,  other  metals  in  the  electrolyte  are 
decomposed,  and  under  aggravated  conditions  the  electrolyte  itself  may 
be  decomposed.  The  relative  amounts  of  the  various  metals  in  solution 
also  determines  the  efficiency  of  the  process  under  certain  conditions. 
If  the  electrolyte  contains  only  a  small  amount  of  copper  and  consideral)le 
zinc — conditions  quite  likely  to  occur — the  copper  will  be  deposited  to 
the  exclusion  of  the  zinc  with  a  correspondingly  low  current  density,  but 
if  that  density  is  exceeded,  the  zinc  will  be  electrolyzed,  and  if  the  elec- 
trolyte is  quite  acid  it  will  be  redissolved  at  the  cathode  as  rapidly  as  de- 
posited, with  the  net  result  that  the  copper  will  be  permanently  deposited 
at  the  cathode,  but  at  a  greatly  increased  expenditure  of  energy.  If 
the  metal  at  the  cathode  is  not  readily  soluble  in  the  electroh'te,  the 
copper  will  be  correspondingly  impure. 

AVhen  copper  ores  are  leached  with  any  acid  solution,  or  a  solution 
having  an  acid  base,  any  or  all  of  the  metals  may  be  in  the  solution,  and 
when  electrolyzed  may  be  influenced  by  the  current  as  described.  As  the 
electrolyte  remains  more  acid,  purer  copper  is  likely  to  be  deposited, 
but  when  it  becomes  neutral,  or  only  slightly  acid,  the  undesirable  metals 
are  deposited  with  the  copper  and  are  likely  to  remain  on  the  cathode. 
Such  a  condition,  however,  is  not  likely  to  occur  since  a  neutral,  or 
only  slightly  acid  solution,  is  not  an  energetic  solvent  of  copper  from 
its  ores. 

The  principal  factors  which  determine  the  kind  of  ion  to  be  deposited 
on  the  cathode  is  the  heat  of  formation  of  the  different  possible  reactions; 
that  reaction  which  absorbs  the  least  energy  occurs,  in  general,  the  most 
readily;  but  this  statement  is  only  approximately  true.  The  concentra- 
tion of  the  two  electrolytes  and  the  current  density  rrtust  be  taken  into 

The  heat  of  formation  of  the  most  important  chlorides  and  sulphates 


likely  to  be  in  the  electrolyte  from  leaching  copper  ores,  in  dilute  solutions, 
is  as  follows: 

Molecular  Heats  of  Formation  of  Chlorides  and  Sulphates  in  Dilute  Aqueous 





96,600  calories 




187,400  calories 




187,100  calories 




128,000  calories 




238,100  calories 




113,000  calories 




100,100  calories 




127,850  calories 


•  650,500 


77,900  calories 




35,400  calories 




62,500  calories 




71,500  calories 




91,400  calories 




90,800  calories 


29,000  calories 


27,200  calories 




39,400  calories 




93,900  calories 



The  decomposition  values  are  independent  of  the  solution,  in  the  case 
of  bases  and  acids  which  on  electrolytic  decomposition  evolve  oxygen  and 
hydrogen  at  the  electrode,  and  this  is  true  for  all  acids  excepting  those 
whose  decomposition  values  are  below  the  maximum.  For  these  the 
value  rises  with  increasing  dilution,  and  finally  reaches  the  maximum. 
This  is  very  marked  in  the  case  of  hydrochloric  acid: 

Decomposition,  point 

2n  HCl,  1 .  26  volts 

n  HCl,  1.31  volts 

l/2nHCl,  1.34  volts 

l/6nHCl,  1.43  volts 

l/16nHCl,  1.62  volts 

l/32nHCl,  1.69  volts 

With  l/32n  HCl  a  point  is  reached  where  chlorine  is  no  longer  given 
off,  but  a  large  proportion  of  oxygen. 

In  leaching  ores  with  sulphate  solutions,  the  chlorides  are  not  likely 
to  present,  but  if  a  chloride  solution  is  used,  the  sulphates  are  quite  sure 
to  be  present,  especially  if  the  ore  is  roasted.  The  chlorides  will  be 
decomposed  before  the  sulphates. 

If,  for  instance,  the  electrolyte  contains  CuCls  ZnClj,  and  NajSO^, 
either  the  Cu,  Zn,  or  Na  may  be  active  in  transporting  the  current 
through  the  solution,  but  the  heat  of  combination  of  cupric  chloride  is 
62,500  calories;  that  of  zinc  chloride  113,000  calories,  and  that  of  sodium 


sulphate  328,500  calories,  so  that  copper  will  be  deposited  to  the  exclusion 
of  zinc  until  the  minimum  voltage  for  the  electrolysis  of  zinc  chloride  is 
exceeded;  then  the  zinc  chloride  may  be  decomposed  if  there  are  not 
enough  copper  ions  in  the  solution  to  transport  the  current  through  the 
electrolyte,  and  finally  even  the  sodium  sulphate  may  be  decomposed,  if 
the  voltage  is  sufficiently  high.  It  is  evident,  however,  even  with  impure 
electrolytes,  copper  may  be  deposited  to  the  exclusion  of  other  metals 
in  appreciable  amounts. 

Effect  of  Bismuth,  Arsenic,  and  Antimony,  in  the  Electrolyte. — None 
of  the  metals  highly  injurious  to  copper,  such  as  bismuth,  arsenic  and 
antimony,  are  likely  to  affect  the  deposited  copper.  These  metals  are 
largely  eliminated  in  the  cyclic  operation  of  the  process,  even  if  contained 
originally  in  the  ore.  If  the  ore  is  a  sulphide  and  has  to  be  roasted,  these 
elements  are  largely  volatilized  during  the  roasting.  If  they  should 
accumulate  in  the  solution,  they  are  easily  removed  and  their  removal 
may  be  made  to  render  an  equivalent  in  acid  by  precipitating  with 
hydrogen  sulphide.  Bismuth  is  not  deposited  on  the  cathode  even  when 
present  in  considerable  quantities.  Neither  arsenic  nor  antimony  are 
deposited  with  the  copper  unless  the  solution  approaches  neutrality. 

Iron  in  the  Electrolyte. — Iron  is  most  likely  to  be  in  the  electrolyte, 
as  it  always  is  associated  with  copper  ores.  In  sulphate  solutions  the 
iron  is  likely  to  accumulate  indefinitely,  to  saturation,  unless  the  solution 
is  purified  at  intervals.  With  chloride  solutions,  the  iron  chloride  acts 
more  or  less  on  the  copper  compounds  in  the  ore  to  form  copper  chloride, 
while  the  iron  is  precipitated  as  the  insoluble  ferric  oxide.  In  the 
electrolysis  of  a  copper  sulphate  solution,  containing  considerable  iron 
sulphate,  the  iron  passes  from  the  ferrous  to  the  ferric  condition  at  the 
anode,  and  is  transformed  back  again  to  the  ferrous  condition  at  the 
cathode,  and  thus  using  energy  without  rendering  a  useful  equivalent. 
Similarly  the  ferrous  chloride  may  be  transformed  to  ferric  chloride  and 
back  again,  as  the  solution  passes  from  anode  to  cathode,  but  in  so  doing 
the  ferrous  chloride  is  likely  to  be  oxidized  to  the  ferric  oxide,  under  the 
oxidizing  influence  of  the  chlorine. 

E.  H.  Larrison,  aptly  sums  up  the  effect  of  iron  sulphate  in  an  electro- 
lyte of  copper  sulphate  as  follows:* 

"The  process  of  electrolysis  exercises  in  the  main  a  reducing  influence  in  an 
electrolyte  of  copper  sulphate  containing  iron  sulphate.  Further  reduction,  and 
probably  the  retardation  of  the  deposition  of  the  copper,  comes  about  through 
some  such  reaction  as  the  following : 

Cu+Fe2(SOj3  =  CuSOi  +  2FeSO,. 

That  is,  copper  already  deposited  or  on  the  point  of  being  deposited  is  attacked 
by  the  ferric  sulphate  and  dissolved  thereby,  also  reducing  the  ferric  to  ferrous 

^E.  andM.J.,  Sept.  7,  1907. 


sulphate.  When  the  process  has  gone  on  for  some  time  the  electrolyte  becomes 
so  dilute  with  respect  to  ferric  sulphate  that  the  last  of  the  copper  is  able  to  keep 
its  place  upon  the  cathode.  This  accounts  for  the  difficulty  of,  and  the  compara- 
tive great  time  necessary  for  the  removal  of  the  last  few  miligrams  of  copper  from 
a  solution  high  in  iron.  When  using  a  stirrer  and  high  current  densities,  a  very 
small  reduction  in  the  current  density  is  quickly  followed  by  the  re-solution  of 
much  of  the  deposited  copper.  It  seems  as  if  the  copper  is  able  to  keep  its  place  on 
the  cathode  only  with  a  certain  electrical  or  reducing  tension. 

"Circulation  of  the  electrolyte  not  only  allows  the  use  of  higher  current 
densities,  and  thereby  more  rapid  work,  but  it  gives  the  current  greater  efficiency. 
This  is  not  only  due  to  the  fact  that  the  solution  is  kept  uniform  and  the  copper 
deposits  more  rapidly,  but  also  the  reducing  effect  of  the  iron  is  much  greater 
and  consequently  the  dissolving  power  of  the  solution  is  decreased  much  sooner. 
A  very  rapid  method  of  reducing  the  iron  before  applying  the  current  would  save 

"The  influence  of  iron  on  copper  electrolysis  may  be  summed  up  as  follows: 

"  1.  High  percentages  of  iron  in  the  ferric  condition  materially  retard  copper 
deposition.     Ferrous  salts  have  little  or  no  effect. 

"2.  The  effect  of  the  electric  current  in  the  usual  solution  is  to  reduce  ferric 
salts  to  ferrous  salts.  This  reduction  must  proceed  to  a  certain  point  before  all 
the  copper  will  remain  on  the  cathode.  This  point  is  about  the  same  whatever 
the  iron  and  copper  percentages,  but  it  varies  some  as  the  solution  is  circulated 
or  stationary. 

".3.  Rapid  circulation  of  the  solution  by  a  stirrer  permits  the  use  of  high 
current  densities  without  sponging,  thereby  making  faster  deposition.  The 
current  also  is  more  efficient  for  both  depositing  copper  and  reducing  iron." 

Purification  of  the  Electrolyte. — Various  methods  have  been  suggested 
for  purifying  the  electrolyte,  and  these  depend  to  a  large  extent  on  the 
nature  of  the  solvent.  Many  are  based  on  the  practice  of  the  electrolytic 

The  solution  may  be  treated  electrolytically  by  passing  a  current  of 
electricity  through  it  with  a  high  current  density  at  the  cathodes,  and 
employing  either  copper  or  lead  anodes  for  sulphate  solutions,  and 
carbon  anodes  for  chloride  solutions.  By  this  means  many  of  the 
metallic  impurities  can  be  thrown  down  on  the  cathode,  especially  if  the 
solution  is  not  too  highly  acid. 

At  Anaconda  a  process  was  used  which  consisted  in  passing  the  impute 
electrolyte  repeatedly  through  a  layer  of  oxidized  copper,  so  as  to  partially 
precipitate  the  antimony  and  bismuth.  By  this  treatment  the  solution 
becomes  nearly  neutral,  and  saturated  with  copper,  and  was  then  oxidized 
by  passing  air  through  it,  so  that  the  iron  is  partially  precipitated  as 
ferric  oxide. 

Ulke  states  that  one  of  the  best  methods  for  purifying  old  solutions  is 
that  in  which  it  is  electrolyzed  in  special  vats,  the  anodes  being  of  lead, 
and  the  cathodes  of  copper.     A  current  density  is  employed  sufficiently 


great  to  deposit  the  arsenic  and  antimony,  but  not  strong  enough  to 
deposit  the  iron.  The  solution  thus  freed  from  the  arsenic  and  antimony 
is  returned  to  the  copper  depositing  vats  to  be  used  in  the  ordinary  way, 
and  this  is  repeated  until  the  bath  contains  so  much  iron  that  it  is 
necessary  to  remove  it  by  crystallizing  out  the  ferrous  sulphate. 

Ottaker  Hofmann  gives  the  following  method  of  purifying  copper 
sulphate  solutions  having  in  addition  to  the  cupric  sulphate,  salts  of  iron, 
arsenic,  antimony,  bismuth,  cobalt,  nickel,  etc' 

"The  crude  neutral  cupric  sulphate  solution  is  forced  into  towers,  about  6  ft. 
in  diameter  and  16  ft.  high.  These  towers  are  lined  with  lead.  A  lead  pipe, 
connecting  with  an  air  compressor,  is  led  into  the  tower  through  the  funnel- 
shaped  bottom.     Near  the  bottom  is  a  lead  steam  coil  for  heating  the  solution. 

"When  the  tower  is  filled  with  impure  solution,  obtained  by  treating  roasted 
copper  matte  with  dilute  sulphuric  acid,  steam  is  allowed  to  enter  the  coil  and 
heat  the  solution  and  at  the  same  time  air  is  forced  through  the  pipe  at  the 
bottom.  The  ascending  air  imparts  a  violent  boiling  motion  to  the  liquor. 
Part  of  the  iron  is  precipitated  as  basic  salt,  by  the  action  of  the  air.  More  than 
half  the  iron  was  never  precipitated  although  the  treatment  was  extended  many 
hours.  When  the  solution  is  hot  (75  to  80°  C.)  roasted  matte  is  added.  The 
violent  boiling  motion  of  the  solution  keeps  the  matter  in  suspension,  and  after 
3  or  4  hours  the  solution  will  be  entirely  free  from  iron,  arsenic,  antimony,  bis- 
muth, etc.  To  observe  and  regulate  the  progress  of  the  operation  the  solution 
is  tested  from  time  to  time  for  iron  by  taking  samples  through  a  small  coc'k 
inserted  in  the  sides  of  the  tower.  It  is  not  necessary  to  test  the  solution  for 
other  impurities,  because  the  iron  predominates,  and  by  the  time  all  of  it  has  been 
precipitated,  no  trace  of  the  other  impurities  will  be  found." 

"The  chemical  reaction  of  this  process  may  be  expressed  as  follows: 

2FeS0i  +  0  +  CuO  =Fe203  +  2CuS0<. 

This  shows  that  the  cupric  oxide,  which  together  with  air,  is  used  to  precipitate 
the  iron,  combines  with  the  sulphuric  acid  of  the  ferrous  sulphate,  and  goes  in 
solution  as  cupric  sulphate;  a  decided  advantage,  as  the  precipitant  is  converted 
into  cupric  sulphate,  and  thus  enriches  the  copper  solution." 

At  the  Kalakent  Copper  Works,  Russia,^  when  the  impurities  in  the  elec- 
trolyte had  accumulated  to  such  an  extent  as  to  endanger  the  quality  of 
the  electrolytic  copper,  the  foul  solutions  were  withdrawn  from  the  in- 
dividual or  group  of  tanks  and  regenerated,  this  regeneration  being  ac- 
complished with  considerable  difficulty  at  first,  but  finally  it  was  done 
advantageously  as  follows: 

"The  foul  solutions  were  heated  and  passed  over  dead  roasted  matte  fines, 
heaped  in  loose-bottomed  trays  arranged  in  series  in  upright  rows.  By  this 
method  all  the  sulphuric  acid  in  solution,  down  to  about  2  grm.  per  100  c.c,  was 
neutralized  and  combined  with  copper  and  iron.     The  solution  was  then  allowed 

'Mineral  Industry,  Vol.  VIII,  p.  192. 

^  Titus  Ulke,  "Modern  Electrolytic  Copper  Refining,"  p.  145. 


to  trickle  through  heaps  of  roasted  low-grade  copper  ores,  which  resulted  in  the 
almost  complete  neutralization  of  its  acid  contents.  It  was  then  diluted  with 
washwaters  down  to  10  to  12°  B.  and  heated,  to  free  the  solution  from. iron, 
arsenic,  tin,  and  bismuth,  in  lead  pans,  which  were  provided  with  an  appliance 
for  injecting  compressed  air  into  the  solution.  Anode  scrap  was  suspended  in 
the  pans  so  as  to  neutralize  any  remaining  portion  of  acid,  and  to  take  up  any- 
new  acid  set  free  through  the  separation  of  iron  hydrates,  and  thereby  form 
copper  sulphate.  By  blowing  compressed  air  into  the  solution  heated  to  about 
50°  C,  the  copper  in  the  anode  residue  or  scrap  was  quickly  dissolved,  and  the 
separation  of  the  iron,  arsenic,  antimony,  tin,  and  bismuth,  which  occurred  when 
the  solution  had  been  nearly  or  completely  neutrahzed,  accomplished.  The 
solution  was  then  concentrated  up  to  14  to  15°  B.  and  clarified  in  special  reser- 
voirs. It  now  contained  3.5  to  4  grm.  of  copper  per  100  c.c,  and  of  impurities 
only  0..5  to  1  grm.  iron,  besides  traces  of  zinc,  nickel,  and  cobalt,  and  was  there- 
fore pure  enough  for  reuse  as  electrolyte.  The  excess  of  purified  electrolyte 
which  gradually  accumulated  with  this  method  of  regeneration  was  eventually 
withdrawn  from  the  circulating  system  and  worked  up  into  blue  stone.  The 
electrolytic  copper  possessed  a  purity  of  at  least  99.9  per  cent,  by  analysis,  and 
averaged  99.93  per  cent,  copper." 

If  cupric  sulphate  is  crystallized  out  of  impure  solutions  and  redis- 
solved  in  water,  it  may  be  electrolyzed  to  deposit  the  copper  and  liberate 
the  combined  sulphuric  acid.  The  acid  solution  may  then  be  again 
applied  to  the  ore,  with  all  or  'most  all  of  the  impurities  eliminated. 

Hoepfner'  proposed  precipitating  the  impurities  from  a  chloride 
solution  with  oxychloride  of  copper.  In  doing  this,  some  of  the  cuprous 
chloride  in  the  solution  is  converted  into  the  oxychloride  of  copper  by 
blowing  air  or  oxygen  into  the  electrolyte: 

CU2Cl2  +  0  =  CU2Cl20. 

"This  is  most  conveniently  done  by  cooling  the  solution,  or  a  part  thereof  to 
precipitate  the  cuprous  chloride,  which  is  then  converted  into  the  oxychloride  by 
contact  with  air.  This  solid  precipitate  is  a  most  efficient  reagent  for  iron  and 
to  enrich  solutions  poor  in  copper  so  as  to  make  it  more  suitable  for  electrolysis. 
The  reaction  may  be  expressed  as : 

3CU2CI.O  +  2FeCl2  =  4CUCI2  +  Cu^Clj  -l-FejOs. 

Lime,  the  caustic  alkalis  or  alkaline  earths,  or  their  carbonates,  may  be  used  as 
precipitants ;  or  the  oxides  or  carbonates  of  metals,  as,  for  instance,  of  copper, 
may  be  employed  in  the  separation  of  the  undesirable  metals  from  the  electrolyte, 
the  metals  being  precipitated  according  to  the  precipitant  used  in  the  form  of 
oxide  or  in  the  form  of  arsenate  and  antimonate  of  iron  or  copper,  the  arsenic 
and  antimony  being  present  in  the  solution  in  the  form  of  arsenious  and  antimon- 
ious  acids,  or  arsenic  and  antimonic  acids  which  are  converted  by  the  precipitant 
into  insoluble  arsenic  and  antimonic  salts,  while  if  an  oxide  or  carbonate  of 
copper  is  used  the  arsenic  and  antimony  are  converted  into  insoluble  arsenate  or 

lU.  S.  Patents  No.  507,130,  Oct.  24,  1893,  and  No.  704,639,  July  15,  1902. 


arscnite  of  copper  and  the  corresponding  salts  of  antimony.  Inasmuch  as  the 
electrolyte  contains  copper  it  may  readily  happen  that  the  salts  last  referred  to 
will  be  formed  without  the  use  of  a  copper  salt  when  lime  or  an  alkali  is  used  as 
the  precipitant.  In  either  case,  but  a  comparatively  small  proportion  of  the 
copper  goes  over  with  the  precipitant,  which  quantity  is  in  no  ease  greater  than 
t!ie  quantity  of  the  arsenic  and  antimonic  salts  precipitated." 

Greenawalt  purposes  purifying  chloride  solutions  by  electrolyzing 
common  salt  to  form  caustic  soda  and  chlorine;  the  caustic  soda  is  used 
as  a  precipitant  of  the  impurities,  and  the  chlorine  to  produce  acid. 

In  purifying  the  electrolyte,  a  certain  portion  is  withdrawn  from  the 
cycle  of  operation,  and  the  caustic  soda  applied  to  it; 

RCl2  +  2NaOH  =  2NaCl  +  R(OH), 

in  which  R  may  represent  any  or  all  the  base  metals. 

In  this  way  the  base  metals  are  precipitated  and  the  sodium  chloride 
regenerated.  The  chlorine  liberated  by  electrolysis  of  the  salt  is  combined 
with  sulphur  dioxide,  in  the  presence  of  water,  or  the  solution,  to  form 

2Cl  +  S02+2H20=2HCl  +  H2SOi. 

In  this  way  all  the  undesirable  impurities  are  eliminated  from  the  solu- 
tion, and  a  corresponding  amount  of  acid  solution  regenerated. 

In  the  purification  of  the  electrolyte  it  is  generally  assumed  that  a 
pure  deposit  of  copper  is  necessary.  If  the  impurities  in  the  electrolyte 
do  not  materially  interfere  with  the  efficiency  of  the  process  it  would 
be  better  to  work  with  impure  solutions  even  though  an  impure  copper 
is  deposited.  There  is  no  reason  why  the  purity  of  the  deposited  copper 
should  be  a  matter  for  serious  consideration. 

The  relative  efficiency  of  the  process,  due  to  impurities  in  the  electro- 
lyte is  more  serious  than  the  relative  purity  of  the  copper.  Some  ele- 
ments cause  serious  loss  of  efficiency  if  present  in  certain  combinations, 
while  in  other  combinations  the  loss  cannot  be  considered  of  much 
consequence.  Ferric  salts,  for  example  in  the  electrolyte,  may  cause  a 
serious  loss  of  efficiency,  while  ferrous  salts  are  comparatively  harmless. 
And  this  applies  generally  to  the  elements  which  have  different  valencies 
for  different  combinations,  and  which  are  capable  of  oxidation  and 
reduction  in  the  electrolyte  due  to  electrolysis. 

Depolarizers. — Various  depolarizers  have  been  suggested  in  connection 
with  the  electrodeposition  of  copper;  among  the  most  important  is  sul- 
phur dioxide.  The  use  of  reducing  gases,  and  particularly  sulphur 
dioxide,  for  the  depolarization  of  insoluble  anodes  in  the  deposition  of 
copper,  was  described  as  long  ago'  as  1878  by  Cobly.  Later  Luckow 
used  sulphur  dioxide  in  the  electrolysis  of  zinc  solutions,  and  its  applica- 
tion in  more  recent  years  has  been  quite  common  and  well  understood. 


If  cupric  sulphate  is  electrolyzed: 

CuSO  4  + electric  current  =  Cu  + SO  4. 

The  resulting  products  are  copper  at  the  cathode  and  sulphion  at  the 
anode.  The  liberation  of  the  SO  4  at  the  anode  immediately  results  in  a 
secondary  reaction, 

S04  +  H20  =  H2S04  +  0 

whereby  water  combines  with  the  sulphion  to  form  sulphuric  acid,  while 
oxygen  is  released.  If,  however,  sulphur  dioxide  is  in  the  anode  solution, 
the  SO4  and  SOj  combine,  with  water,  to  form  two  atoms  of  sulphuric 

S04  +  S02  +  2HjO  =  2H2S04 

and,  theoretically,  no  oxygen  is  released.  This  reaction  develops  a 
certain  electromotive  force  at  the  anode,  in  the  same  direction  as  the 
current,  and  thus  reduces  the  necessary  voltage  in  the  decomposition  of 
the  copper. 

If  sulphur  dioxide  is  used  in  connection  with  chloride  solutions,  the 
electric  current  decomposes  the  copper  chloride  into  copper  and  chlorine, 

2CuCl  + electric  current  =  2Cu  +  2C1. 

The  chlorine  combining  with  the  water, 

2C1  +  H20  =  2HC1 +  0  +  10,800  calories, 

to  produce  hydrochloric  acid  and  release  oxygen,  and  thus  chlorine  may 
become  an  oxidizing  agent.  This  reaction  takes  place  very  slowly,  but 
the  oxygen  can  immediately  exert  a  further  chemical  action  in  the 
presence  of  sulphur  dioxide,  when  a  rapid  decomposition  of  water  takes 

2Cl  +  SO2  +  2H2O  =  H2SO4  +  2HCl  +  74,400  calories 

in  which  case  the  chlorine  released  at  the  anode  is  by  secondary  reaction, 
converted  into  sulphuric  and  hydrochloric  acids  with  the  development  of 
74,400  calories,  which  tends  to  reduce  the  voltage  in  the  operation. 
Ordinarily,  chlorine  will  be  given  off  at  the  anode  in  the  absence  of  sulphur 
dioxide,  because  the  oxidizing  action  of  chlorine  is  exceedingly  slow;  in 
the  presence  of  sulphur  dioxide  the  amount  of  chlorine  released  will 
depend  upon  the  current  density.  If  the  current  density  is  sufficiently 
low  to  permit  of  the  released  chlorine  coming  in  contact  with  the  sulphur 
dioxide  before  either  can  escape,  no  chlorine  will  appear,  as  such. 

Other  reactions  have  been  suggested  in  the  decomposition  of  copper 
compounds,  acting  on  the  principle  of  depolarizers.  In  Body's  process, 
ferrous  chloride  is  converted  into  ferric  chloride  by  the  chlorine  released 


at  the  anode;  in  the  Siemens-Halske  process  ferrous  sulphate  is  recon- 
verted to  the  ferric  sulphate  in  the  deposition  of  the  copper  from  a  solu- 
tion of  cupric  and  ferrous  sulphates;  and  in  the  Hoepfner  process,  in 
which  the  copper  is  deposited  from  a  cuprous  chloride  solution  and  a 
certain  amount  of  the  cuprous  chloride  converted  back  to  the  cupric 
chloride  by  the  chlorine  released  at  the  anode. 

In  all  of  these  reactions,  the  electromotive  force  developed  works  in 
the  direction  of  the  current  and  thus  reduces  the  necessary  voltage.  The 
approximation  of  the  theoretical  to  that  realized  in  practice,  depends  on 
the  conditions  of  operation,  principally  among  these  conditions  is  the 
current  density.  If  the  current  density  is  so  low  as  to  permit  all  of  the 
anode  gases  to  combine  with  the  depolarizer,  the  theoretical  efficiency 
may  be  quite  closely  realized. 

Rapid  Deposition  of  Copper. — To  rapidly  deposit  a  metal  from  solution 
by  the  electric  current,  it  is  necessary  that  the  metal  ions  be  present  in 
sufficient  number  at  the  cathode.  If  a  comparatively  high  cathode  density 
is  used,  the  danger  is  that  the  electrolyte  in  proximity  to  the  cathode 
becomes  poor  in  the  ions  deposited,  and  other  processes  start,  especially 
the  development  of  hydrogen.  To  counter  this  tendency,  it  is  necessary 
to  artificially  bring  fresh  electrolyte  to  the  cathode  surface,  as  for  instance, 
by  stirring,  or  rotating  the  cathode. 

In  an  ordinary  electrolyzer,  if  the  electrolyte  is  not  agitated  or  cathode 
rotated,  as  soon  as  the  current  is  switched  on,  copper  is  deposited  on  the 
cathode  and  a  thin  layer  of  electrolyte  touching  the  cathode  becomes  in 
consequence  impoverished  in  copper.  Before  more  deposition  can  take 
place  this  thin  layer  of  exhausted  electrolyte  has  to  be  removed.  If  left 
to  itself,  removal  and  changing  will  take  place  quite  slowly,  nevertheless 
it  will  be  fast  enough  to  supply  sufficient  copper  ions  to  the  cathode  if 
the  current  density  is  correspondingly  small.  No  loss  of  efficiency  would 
therefore  result.  If,  however,  the  current  be  greater  than  the  correspond- 
ing removal  of  the  impoverished  electrolyte  from  the  cathode,  the  current 
will  sieze  upon  the  next  available  material  in  the  electrolyte,  and  the 
result  will  be  a  corresponding  loss  of  efficiency  and  impure  and  rough 
copper  deposited.  To  prevent  this  action  from  taking  place  it  is  neces- 
sary to  remove  the  film  of  exhausted  electrolyte  as  rapidly  as  it  is  formed, 
either  by  violent  circulation  of  the  electrolyte,  friction,  or  rotating  cath- 
ode. Consequently  the  process  which  most  effectually  removes  the 
impoverished  electrolyte  from  the  cathode,  and  which  at  the  same  time 
will  give  the  most  uniform  distribution  of  current  and  the  most  uniform 
strength  of  electrolyte  over  the  whole  surface  of  the  cathode,  should  be 
the  one  to  produce  the  highest  efficiency,  the  best  deposits  of  copper,  and 
allow  of  the  highest  current  density. 

Various  methods  are  employed  to  accomplish  these  results.  One 
generally  used,  is  to  place  the  electrolyzers  in  cascade  series,  that  is  to 




say,  in  which  there  is  a  slight  difference  in  level  between  one  electrolyzer 
and  the  one  next  to  it,  so  that  the  electrolyte  pumped  into  a  tank  above 
the  highest  electrolyzer  is  fed  into  it,  and  gradually  flows  through  the 
entire  series  and  finally  issues  from  the  lowest  in  the  series  and  is  again 
pumped  back  to  the  tank,  or  to  the  leaching  vats.  The  rapidity  of  the 
circulation  is  governed  by  the  difference  in  level  between  any  two  tanks 
and  the  slant  of  the  tank  itself.  If  the  tank  has  not  sufficient  slant,  the 
electrolyte  will  overflow  if  the  circulation  is  at  all  rapid. 

Fig.  51. — Coffin  revolving  cathode  apparatus. 

Another  method  of  accomplishing  the  same  purpose  is  to  force  air  into 
the  tank  and  electrolyte  while  the  electrolysis  is  in  progress.  Care 
should  be  used  if  the  electrolyte  is  agitated  with  air,  that  the  agitation  is 
quite  uniform,  in  all  parts  of  the  electrolyzer.  Sometimes  stirrers, 
working  reciprocally  between  the  electrodes  are  used  to  better  advantage 
than  air,  and  the  agitation  is  likely  to  be  more  uniform.  The  electrolyte, 
impinging  against  the  cathodes  in  jets  has  given  good  results,  but  is  now 









Fig.  553. — Hoepf ner  revol  ving  cathode 
apparatus     Section. 

nowhere  used.  A  rotating  cathode  gives  excellent  results,  but  a  very 
different  type  of  cell  is  required  than  when  any  of  the  other  methods  are 
employed.  If  diaphrams  must  be  used,  a  rotating  cathode  may  offer 
some  constructional  difficulties,  which,  however,  need  not  be  insur- 
mountable. Fig.  54  shows  one  form  of  revolving  cathode  apparatus 
devised  by  Coffin'  in  which  A  is  the  electrolyzer,  lined,  preferably,  with 
sheet  lead,  fi  is  a  post  or  pedestal  secured  in  the  center  of  the  tank  and 
surmounted  by  a  metallic  head  D,  or  socket,  E  is  the  cathode  cylinder, 
and  0  the  anodes.     Figs.  55  and  55i  represent  another  form  of  apparatus 

devised  by  Hoepfner'  in  which  d  indi- 
cates the  carbon  anodes,  D  the  dia- 
phragms, and  a  the  cathodes,  revolving 
on  a  horizontal  shaft  c.  In  addition 
to  the  revolving  cathode,  brushes  are 
sometimes  used  and  adjusted  so  as  to 
cause  friction  on  the  deposited  cathode 
material.  Cowper-Coles  found  that 
copper  possessing  the  advantages  of 
hard  rolled  copper  of  high  tensile 
strength  and  free  from  porosity  can  be 
deposited  to  any  thickness  desired  at ' 
a  rapid  rate  by  revolving  the  cathode 
at  a  peripheral  speed  of  from  1500  to  2000  ft.  per  minute  when  employ- 
ing current  densities  of  200  amperes  per  square  foot  of  cathode  surface 
and  an  electrolyte  containing  12.5  per  cent,  of  copper  sulphate  and  13 
per  cent,  of  sulphuric  acid  at  a  temperature  of  40°  C 

In  experimental  work  in  the  laboratory  of  applied  chemistry  at  the 
University  of  Wisconsin''  it  was  observed  that  the  critical  current  density, 
that  is  to  say  the  current  density  at  which  powdery  deposit  occurs,  is 
approximately  proportional  to  the  speed  of  rotation,  or  better,  to  the 
linear  feet  per  minute,  as  the  cathode  speed  depends  upon  the  speed  of 
rotation  and  the  diameter  of  the  revolving  cathode. 

Betts^  uses  anodes  preferably  in  the  form  of  rods,  and  gives  to  them 
a  reciprocating   motion  in  a  direction  perpendicular  to  their  length, 
whereby  the  layer  of  electrolyte  touching  the  anode  is  rapidly  changed. 
In  electrolyzing  a  solution  of  ferrous  and  cupric  sulphates,  with  a  dia- 
phragm, depositing  copper  on  a  cathode,  and  converting  ferrous  to  ferric 
sulphate  at  the  anode,  and  with  a  circulation  of  electrolyte  that  was  pre- 
viously considered  amply  sufficient,  the  electromotive  force  required  to 
work  the  cell  was  considerably  reduced,  the  evolution  of  gas  at  the 
'  U.  S.  Pat.  415,024,  Nov.  12,  1889. 
=  U.  S.  Pat.  598,180,  Feb.  1,  1898. 
'  U.  S.  Patent  895,163,  Aug.  4,  1908. 

*  J.  G.  Zimmerman,  N.  Y.  Meeting  Electrochemical  Society,  1904. 
=  U.  S.  Pat.  803,543  Nov.  7,  1905. 


anode  was  entirely  stopped,  and  the  current  efficiency  raised  from  about 
50  per  cent,  to  100  per  cent,  on  giving  the  anodes  a  reciprocating  motion 
of  about  one  hundred  complete  cycles  per  minute,  with  an  amplitude  of 
about  1  in. 


The  electrolytic  sulphate  processes  may  be  divided  into  two  general 
classes,  based  on  the  solvent,  as, 

Sulphuric  Acid, 
Ferric  Sulphate, 

but  it  is  evident  that  neither  of  these  processes  can  be  carried  out  to  the 
exclusion  of  the  other.  All  solutions  resulting  from  leaching  roasted  or 
oxidized  ores  will  have  more  or  less  iron  sulphate  either  in  the  ferric  or 
ferrous  condition,  and  the  utility  of  a  neutral  solution  of  ferric  sulphate, 
as  compared  with  an  acid  solution,  is  questionable.  Sulphuric  acid  is  an 
energetic  solvent  of  copper  from  roasted  or  oxidized  ores,  and  is  to  be 
generally  preferred.  On  certain  chalcocite  ores  ferric  sulphate  has  given 
good  results  without  roasting. 

In  a  copper  sulphate  solution  containing  ferrous  sulphate,  both 
•  sulphuric  acid  and  ferric  sulphate  may  be  regenerated  during  electrolysis, 
and  to  the  formation  of  ferric  sulphate  much  of  the  difficulties  in  the  elec- 
trolysis of  impure  copper  sulphate  solutions  may  be  attributed.  If  a  solu- 
tion of  copper  sulphate,  as  for  example  that  derived  from  leaching  copper 
ores,  is  electrolyzed  without  a  diaphragm,  then  in  addition  to  the  reaction; 

CuSO  4  +  H^O  +  Electric  C  '.rrent  =  Cu  +  H^SO  ^  +  0, 

there  may  also  take  place, 

CuSO 4  +  2FeS0 4  +  Electric  Current  =  Cu  +  Fe2 (SO 4)  3, 
3FeS0 4  +  Electric  Current  =Fe  +re2(S0 4)  3. 

The  ferric  sulphate,  finding  its  way  back  to  the  cathode,  and  under  the 
influence  of  the  current,  gives  rise  to  reversible  reactions,  thus: 

Fe2(S04)3-^Cu  =2FeS04  +  CuS04, 
Fe2(S04)3+Fe  =3FeS04, 

thus  nullifying  the  previous  reactions,  and  resulting  in  a  loss  of  effi- 
ciency. This  loss  will  depend  largely  on  the  amount  of  iron  in  the  solu- 
tion. If  the  iron  is  excessive,  the  copper  may  be  dissolved  as  rapidly  as 
precipitated,  and  the  sum  total  of  the  energy  expended  will  be  nil,  so 
far  as  any  useful  effect  is  concerned. 

If  a  suitable  diaphragm  is  interposed  between  the  electrodes,  then 
both  sulphuric  acid  and  ferric  sulphate  are  regenerated,  but  the  deleterious 
reactions  at  the  cathode  are  avoided,  and  the  current  efficiency  may 
very  closely  approximate  the  theoretical,  but  the  energy  efficiency  will 
be  governed  somewhat  by  the  resistance  of  the  diaphragm.   The  relative 


amounts  of  sulphuric  acid  and  ferric  sulphate  regenerated  at  the  anode, 
will  depend  mostly  on  the  relative  proportions  in  the  solution.  If  the 
ferrous  sulphate  is  highly  concentrated  then  ferric  sulphate  will  be 
largely  regenerated,  and  the  process  becomes  one  in  which  ferric  sulphate 
is  used  as  the  lixiviant  and  in  which  the  resulting  ferrous  sulphate  is  re- 
generated to  the  ferric  sulphate. 

Sulphuric  Acid  Process. — The  simplest  form  of  the  electrolytic  sul- 
phate process  is  when  sulphuric  acid  is  used  as  the  initial  solvent  to 
extract  the  copper  from  its  ores  and  the  sulphate  solution  thus  obtained 
electrolyzed  to  deposit  the  copper  and  again  liberate  the  combined  acid. 
Copper  in  its  sulphide  combinations  is  not  soluble  in  sulphuric  acid;  it  is 
therefore  necessary  to  roast  the  ore  if  a  sulphide  before  the  solution  of 
the  copper  can  be  effected.  After  roasting,  the  copper  will  usually  be  in 
the  form  of  oxide  or  sulphate,  depending  upon  the  nature  of  the  roast. 
In  any  event,  the  sulphuric  acid  may  be  regarded  as  acting  on  the  oxide 
of  copper: 

CuO  +  H^SO,  =  CuSO,  +  H^O. 

The  copper  sulphate  so  produced,  in  addition  to  that  soluble  in  the 
ore,  is  then  electrolyzed: 

CuSO  4  + Electric   current  =Cu  +  S04. 
SO, +  11,0  =  11,80, +  0. 

The  current  decomposes  the  copper  sulphate,  depositing  the  copper  on 
the  cathode,  and  the  acid  radical  is  liberated  at  the  anode.  The  acid 
radical,  combining  with  water,  is  converted  into  an  equivalent  amount 
of  sulphuric  acid  to  that  from  which  the  copper  was  deposited,  and  oxygen 
makes  its  appearance  at  the  anode  as  a  result  of  the  secondary  reaction. 
For  every  pound  of  copper  deposited  1 . 54  lb.  of  acid  is  regenerated. 

If  there  were  nothing  to  consider  but  the  solution  of  the  copper  from 
the  ore  with  sulphuric  acid,  and  the  electrolytic  decomposition  of  the 
resulting  copper  sulphate  solution  into  metallic  copper  and  the  acid 
radical,  this  would  be  a  perfect  process,  requiring  nothing  more  than  the 
expenditure  of  a  certain  amount  of  energy  to  carry  it  on  indefinitely. 
The  difficulties,  however,  in  practically  carrying  out  this  simple  process 
are  considerable. 

Much  of  the  acid  consumed  in  treating  copper  ores  reacts  with  the 
base  elements,  and  hence  there  is  not  sufficient  acid  regenerated  to  treat 
the  next  charge  of  ore;  so  that  the  deficiency  has  to  be  made  up  in  some 
other  way.  The  difficulty,  therefore,  of  the  solution  of  the,  copper  is 
only  partially  solved,  nevertheless  it  is  a  great  step  in  advance  of  iron 
precipitation,  where  all  the  acid  is  irrecoverably  lost.  If  the  ore  is  a 
sulphide,  and  has  to  be  roasted,  some  of  this  difficulty  may  be  overcome 
by  roasting  as  much  of  the  copper  as  possible  to  sulphate  rather  than  to 
oxide.     The  sulphate  of  copper  thus  formed  is  the  same  as  that  produced 


by  the  action  of  sulphuric  acid  on  copper  oxide,  so  that  while  less  acid  is 
used  in  dissolving  the  copper  from  the  ore  more  acid  is  regenerated  in 
the  electrolysis  by  the  amount  of  copper  soluble  in  the  ore  as  sulphate. 
The  deficiency  may  also  be  supplied  by  installing  a  small  acid  works 
in  coimection  with  the  other  metallurgical  operations,  or  it  may  be  pur- 
chased from  acid  manufacturers,  but  both  of  these  methods  add  con- 
siderable to  the  expense,  and  hence  do  not  ingeniously  solve  the  difficulty. 
Further,  the  impurities  in  the  ore,  which  cause  an  irrecoverable  loss 
of  acid,  also  contaminate  the  electrolyte,  and  cause  difficulty  in  the 

No  really  satisfactory  anode  for  sulphate  solutions  has  yet  been 
discovered.  Lead  is  ordinarily  used  for  the  insoluble  anode  in  depositing 
copper  from  sulphate  solutions,  and  while  lead  makes  the  best  anode  for 
sulphate  solutions,  it  is  far  from  being  satisfactory.  In  depositing 
copper  from  sulphate  solutions,  oxygen  is  released  at  the  anode;  this 
oxygen  is  not  entirely  harmless,  but  attacks  the  lead  and  converts  it 
into  the  peroxide,  PbOj.  The  peroxide,  in  addition  to  destroying  the 
anode,  offers  considerable  resistance  to  the  electric  current,  and  thus 
necessitates  an  excessive  consumption  of  power. 

In  a  very  careful  test  made  by  Thomas  P.  Hughes,  in  Denver,  to 
determine  the  peroxidation  of  antimonial  lead  anodes  in  electrolyzing 
copper  sulphate  solution  produced  by  leaching  Arizona  carbonate  ore, 
the  following  results  were  recorded: 

Duration  of  test,  75 .  78  hours. 

Copper  deposited,  6 .  56  lb. 

Average  current,  30.0    amperes. 

Anode  area,  3 . 5    sq.  ft. 

Cathode  area,  3.5    sq.  ft. 

Current  density,  8 . 5    amperes  per  square  foot. 

Average  voltage,  2 . 0    volts. 

Watts,  60.0 

Kilowatt-hours,  4 .  547 

Copper  per  k.  w.-hour,  1 . 4  lb. 

There  were  15  sheet  copper  cathodes,  and  14  antimonial  lead  anodes, 
each  6X6  in.  The  anodes  increased  14  oz.  in  weight.  There  was  a 
black  coating  of  peroxide  of  lead  on  the  anodes,  which  could  easily  be 
scraped  off.  The  increased  weight  of  the  anodes  was  evidently  due  to 
the  oxygen  combined  with  the  lead  to  form  the  peroxide.  As  86.6  per 
cent,  of  peroxide  of  lead  is  lead,  and  13.4  per  cent,  oxygen,  it  follows  that 
5«6  lb.  of  lead  was  peroxidized  in  depositing  6.56  lb.  of  copper;  or  for 
every  pound  of  copper  deposited,  0.85  lb.  of  lead  was  peroxidized.  The 
peroxide  of  lead,  unless  closely  watched,  is  likely  to  drop  to  the  bottom 
of  the  electrolyzer  and  short  circuit  the  current.  In  this  experiment 
the  current  was  not  operating  continuously;  it  was  turned  on  in  the 
morning  and  shut  down  in  the  evening. 


In  experiments  made  by  Greenawalt  to  get  more  information  on  the 
rate  of  oxidation  of  lead  anodes,  one  test  of  1000  ampere-hours,  continu- 
ous run,  with  a  current  density  of  20  amperes  per  square  foot  resulted 
in  depositing  37  oz.  of  copper  and  in  producing  14  oz.  of  lead  peroxide. 
Another  run  of  500  ampere-hours  was  made  in  which  20  oz.  of  copper 
was  deposited  and  3  oz.  of  peroxide  recovered.  While  still  another  run 
of  1458  ampere-hours,  with  a  higher  current  density  produced  60  oz.  of 
copper  and  only  3  1/4  oz.  of  peroxide  of  lead. 

There  is  no  difficulty  in  collecting  the  peroxide  and  again  reducing  it  to 
metallic  lead  to  be  reused  for  anodes,  but  it  necessitates  an  extra  expense 
in  the  operation  of  a  plant  which,  however,  is  largely  compensated  for  by 
the  fact  that  with  the  exception  of  a  small  loss  in  reduction  there  is  no 
cost  for  material  for  replacements. 

The  suggestion  naturally  occurs  that  if  the  peroxide  of  lead  is  the 
ultimate  product  of  the  lead  anode,  why  not  make  an  anode  of  peroxide, 
and  thus  overcome  entirely  the  difficulties  due  to  oxidation?  This  has 
been  tried  repeatedly,  and  Hughes  tried  it  in  various  ways,  but  the 
resistance  to  the  electric  current,  even  with  a  good  conducting  skeleton, 
was  so  high  as  to  put  it  beyond  further  consideration. 

If  sulphur  dioxide  is  used  as  a  depolarizer  in  the  electrolysis  of  copper 
sulphate  solutions,  then,  theoretically,  twice  the  amount  of  acid  com- 
bined with  the  copper  is  regenerated: 

S04  +  S02-h2H20=2H2SO, +  21,320  calories, 

but  it  is  difficult  to  carry  this  out  in  practice,  for  the  reason  that  the 
sulphur  dioxide  and  the  acid  radical,  at  the  moment  of  liberation,  cannot 
be  Ijrought  into  sufficiently  intimate  contact  to  make  it  effective,  espe- 
cially if  a  reasonably  large  current  density  is  used. 

If  sulphur  dioxide  is  used  as  a  depolarizer,  some  of  the  lead  in  the 
anode  will  be  converted  into  the  sulphate,  but  the  action  of  the  forma- 
tion of  sulphate  of  lead  is  very  much  slower  than  in  the  ordinary  elec- 
trolysis where  the  lead  is  converted  into  the  peroxide.  The  lead  sulphate, 
however,  is  more  difficult  to  reconvert  back  into  metallic  lead  than 
the  oxide. 

The  energy  required  to  decompose  an  aqueous  solution  of  copper 
sulphate  is  theoretically,  1.22  volts.  In  practice  it  will  usually  vary 
between  1.5  and  3  volts,  depending  principally  upon  the  current  density 
used.  The  theoretical  output  of  copper  is  therefore,  38.35  lb.  per  h.  p.- 
day,  or  51.43  lb.  per  k.  w.-day,  of  24  hours.  If  a  depolarizer  is  used,  such 
for  example  as  sulphur  dioxide,  then  the  acid  radical  combining  with 
the  sulphur  dioxide  to  form  sulphuric  acid,  will  develop  an  electromotive 
force  working  with  the  current,  and  thus  reduce  the  theoretical  voltage, 
but  the  amount  of  this  reduction  is  limited  largely  by  the  current  density 
employed,  and  the  completeness  with  which  the  two  substances  are 


brought  in  contact  witli  one  another  at  the  moment  the  acid  radical  is 
released  at  the  anode.  Tossizza'  ascertained  by  experiment  that  the 
transformation  of  the  sulphur  dioxide  into  sulphuric  acid  at  the  anode 
gives  rise  to  an  electromotive  force  which  diminishes  the  necessary  volt- 
age and  lowers  it  to  0.2  volt. 

While  the  theoretical  voltage  gives  much  desirable  information,  it  is 
always  best,  indeed  necessary,  to  get  the  voltage  by  direct  experiment 
because  there  are  so  many  factors  which  occur  in  practice  that  do  not 
occur,  and  cannot  be  taken  into  account  in  the  theoretical  determina- 
tions. The  practical  voltage,  once  determined,  is  always  constant,  and 
is  independent  of  the  magnitude  of  the  operation.  This  is  a  factor  which 
can  be  just  about  as  accurately  determined  on  a  small  scale,  on  a  labo- 
ratory basis,  as  in  a  large  working  plant.  Similarly,  the  amount  of  copper 
deposited  from  a  certain  electrolyte,  on  a  small  scale  under  the  conditions 
obtaining  in  practice,  will  always  be  constant,  no  matter  what  the  scale 
of  operations  may  be.  For  this  reason,  the  factor  of  efficiency,  and  of 
voltage,  can  be  quite  accurately  determined  beforehand  for  any  par- 
ticular process  and  made  the  basis  of  calculations  of  a  large  plant  of  any 
sized  unit. 

For  sulphate  solutions,  in  practice,  with  a  current  density  of  5  to  10 
amperes  per  square  foot,  the  voltage  will  usually  be  found  to  vary  from 
1.5  to  2.5  volts.  The  practical  energy  efficiency  can  only  be  determined 
by  weighing  the  copper  deposited  in  a  certain  definite  time,  under  an 
observed  voltage  and  current.  It  will  be  found,  in  making  such  tests, 
that  frequently  the  results  will  be  far  from  the  theoretical,  and  that  the 
temperature,  purity  of  the  electrolyte,  and  current  density  have  much 
to  do  with  the  efficiency.  The  best  that  can  be  done  in  practice  is  to 
approach  the  theoretical  efficiency  although  it  does  not  follow  that 
the  most  efficient  process  in  the  electrolysis  is  the  most  economical  in 

Electrolytic  Extraction  of  Copper  From  Ore  at  Medzianka,  Poland. — 
At  Medzianka,  a  copper  mining  locality  in  Russian  Poland,  about  50 
miles  from  Cacrow,  and  140  miles  east  of  Breslau,  explorations  by  the 
Laszczynski  brothers,  led  to  the  discovery  of  ore  bearing  limestone, 
about  11/4  miles  long  and  150  ft.  wide,  containing  copper  ore  inter- 
spersed in  strips  1/2  in.  thick,  with  calc  spar  and  some  quartz.  The 
mineral  is  almost  entirely  copper  glance,  but  mixed  with  it  is  a  little 
azurite  and  malachite. 

The  produce  of  the  mine  has  been  divided  into  ore  with  50  per  cent,  of 
copper  which  is  separated  underground,  and  mixed  ore  with  16  to  20 
per  cent,  copper  containing  calcite  and  pieces  of  limestone,  which  is 
improved  by  hand  picking,  at  the  surface.     The  ore  as  brought  from  the 

'U.  S.  Pat.  710,346,  Sept.  30,  1902. 


mine  is  crushed  in  rolls,  mixed  with  5  per  cent,  of  damp  brick  earth  and 
moulded  into  blocks,  which  when  dried  by  the  waste  heat  of  the  furnace, 
are  subjected  to  a  partial  roasting  in  a  kiln  fired  from  the  outside, 
with  free  access  of  air,  which  converts  the  copper  into  sulphate  and 

The  roasted  blocks  are  then  crushed  fine  and  leached  in  lead  lined 
wooden  tanks,  with  the  electrolyzed  solution  from  the  electrolytic  cells 
containing  about  5  per  cent,  of  free  sulphuric  acid.  A  liquor  containing 
about  5  per  cent,  of  copper  and  1  per  cent,  free  sulphuric  acid  is  obtained. 
This  solution  is  passed  through  a  filter  press,  to  thoroughly  clarify  it, 
and  then  electrolyzed  in  tanks  of  about  35  cu.  ft.  capacity.  Insoluble 
anodes  of  lead  plates  enclosed  in  cloth  bags,  and  thin  copper  cathodes 
are  used.  A  current  of  1000  amperes  at  2.5  volts,  corresponding  to  a 
current  density  of  about  one  ampere  per  square  decimeter  of  cathode 
surface  (10  amperes  per  square  foot)  is  used,  producing  metallic  copper 
free  from  sulphuric  acid  or  oxygen.  The  deposited  copper,  about  1.1 
grm.  per  ampere-hour,  is  nearly  equal  to  the  theoretical  amount.  The 
power  consumed  per  kilogram  of  copper  is  2.28  k.  w. -hours  or  3  1/2  h.  p. 
(1.3  k.  w. -hours  or  1.6  h.  p. -hours  per  pound  of  copper).  Between  the 
anodes  and  cathodes  there  are  wooden  stirrers,  which  agitate  the  solution 
during  the  entire  electrolytic  process.  The  liquor  is  exhausted  in  from 
36  to  40  hours,  and  then  again  containing  about  1.  per  cent,  copper,  and 
from  1  to  7  per  cent,  free  acid,  is  applied  to  a  fresh  lot  of  ore.  The 
cathodes  remain  in  the  bath  for  about  a  month,  when  the  deposit,  from 
1  to  1  1/4  in.  thick,  is  removed  and  sold.  It  is  of  greater  purity  than  the 
ordinary  electrolytically  refined  copper.  The  four  baths  used  in  the 
process  are  served  by  a  Siemens  dynamo  of  1000  amperes  at  12  volts. 
The  daily  output  of  copper  is  from  225  to  500  lb.  The  entire  process  is 
supervised  by  one  man  in  the  mill  without  any  other  trained  assistance. 

A  vital  point  in  the  success  of  the  process  is  in  the  employment  of 
closely  fitted  bags  or  envelopes  of  thick  cotton  duck  for  the  lead  anodes. 
The  bags,  soaked  with  sulphuric  acid,  exclude  the  iron  salts  and  thus 
overcome  much  of  the  difficulty  from  that  source.  Its  function  is  about 
the  same  as  a  diaphragm.  These  cotton  bags  are  renewed  about  once  a 

The  Laszcynski  process,  used  at  Medzianka,  Russia,  is  described  by 
the  inventor  as  follows:^ 

"If  a  solution  of  sulphate  of  iron,  FeSO^,  is  electrolyzed,  at  the  cathode  the 
bivalent  ferro-ion  is  metallically  deposited  at  the  same  time  the  said  cathion 
comes  in  contact  with  the  insoluble  anode,  there  being  no  diaphragm,  and  is  there 
oxidized  into  trivalent  ferri-ion.  The  latter,  however,  before  it  is  deposited  as 
metallic  iron  has  to  be  reduced  at  the  cathode  to  ferro-ion.  In  this  manner 
there  is  soon  set  up  a  state  of  equilibrium  in  which  the  same  quantity  of  ferro- 

'  U.  S.  Pat.  No.  757,817,  April  1,  19,  1904. 


3  If) 

ions  arc  reduced  at  the  cathode  as  are  produced  at  the  anode.     The  chemical 
action  of  the  current,  therefore,  is  nil." 

"In  the  present  process  the  detrimental  side  action  is  avoided  by  wrapping 
around  the  insoluble  anode  a  cover  or  envelope  of  porous  fabric.  The  wrapping 
being  permeable,  there  is  before  closing  the  circuit  no  difference  between  the 
chemical  composition  of  the  anode  and  cathode  bath.  As  soon  as  the  current 
is  turned  on,  however,  a  layer  of  pure  sulphuric  acid  will  form  around  the 
anode,  since  there  the  SO^  ions  are  discharged.  Consequently  new  sulphuric 
acid  is  generated  which  can  only  drain  off  into  the  close-fitting  envelope,  dis- 
placing in  this  manner  the  solution  of  iron  sulphate,  so  that  in  a  sliort  time 
a  second  process  takes  place  similar  to  the  one  described.  Since  the  ferro-ions 
and  the  ferri-ions  are  cathions,  they  travel  at  the  closing  of  the  circuit  from 
the  anode  to  cathode.  The  envelope  aro>ind  the  anode  forms  a  layer  of  quiet 
liquid,  no  matter  if  the  electrolyte  is  in  circulation,  so  that  the  traveling  of  the 
cathions  can  take  place  without  being  dis- 
turbed. The  result  of  the  two  actions  is 
that  no  ferro-ions  can  be  oxidized,  since 
none  come  in  contact  with  the  anode." 

"Referring  to  the  drawing,  Fig.  56,  in 
the  application  of  the  process  to  cojjper 
ores,  a  is  the  electrolytic  cell,  preferably 
made  of  wood  and  tightened  with  asphaltum 
or  the  like,  b  represents  the  copper  sheet 
cathodes,  and  c  represents  the  anode  which 
consists  of  refined  lead  and  is  provided  with  an  envelope  of  thick  cotton  stuff; 
for  example,  fustian." 

All  copper  ores  without  exception  contain  iron,  which  when  treated  with 
sulphuric  acid  dissolves,  together  with  the  copper.  By  the  electrolysis  the  iron 
is  oxidized  at  the  anode  to  ferric  sulphate,  which  salt  dissolves  the  copper 
deposited  on  the  cathode  equal  to  the  action  of  dilute  nitric  acid : 

Cu-l-Fe,(SO,)3  =  CuSO,  +  2FeSO,. 

In  this  way  the  amount  of  copper  deposited  is  not  only  reduced  to  one-half  or 
even  less,  but  also  a  brittle  and  inferior  metal  is  obtained. 

"The  present  invention  now  prevents  the  oxidation  of  the  iron  salts  and 
makes  possible  the  direct  electrolysis  of  copper  baths  containing  iron,  even  if 
they  contain  twice  as  much  iron  as  copper,  with  a  useful  effect  differing  but 
slightly  from  the  theoretical,  because  the  iron  remaining  in  the  state  of 
wholly  inoffensive  ferrous  sulphate,  FeSO^,  there  is  no  corroding  action  of 
any  kind." 

Fig.  56. 

The  Laszcynski  process  for  electrolytically  obtaining  metals,  especially 
copper  and  zinc,  out  of  the  ores  by  means  of  insoluble  anodes,  consists  in 
tightly  wrapping  the  insoluble  anode  in  a  porous  and  perfectly  permeable 
envelope,  of  fabric  or  other  material,  the  thickness  of  which  is  in  inverse 
proportion  to  the  applied  density  of  current,  for  the  purpose  of  preventing 
anodic  oxidation  of  the  cathions. 


Plant  of  the  Intercolonial  Copper  Co.,  N.  S.  Canada. — The  plant  of  the 
Intercolonial  Copper  Co.,  in  N.  S.  Canada,  was  designed  by  Henry  Car- 
michael,  and  according  to  Johnson'  this  plant  for  some  time  produced 
one  ton  of  electrolytic  copper  daily,  which  was  sold  to  brass  founders 
as  equal  to  the  best  brands  of  electrolytic  copper. 

The  copper  in  the  ore  varied  from  2  to  4  per  cent.  The  very  small 
values  in  the  precious  metals  were  lost.  The  ore  was  crushed  to  20 
mesh,  which  was  fed  by  screw  conveyors  into  a  battery  of  15  revolving 
roasters.  These  consisted  of  long  tubes  of  cast  iron  passing  through  a 
firebrick  muffle,  heated  by  a  flame  from  the  fiire-box.  The  first  part  of 
the  roaster  was  a  brick  lined  drum.  All  revolved,  and  the  ore  was  first 
partially  desulphurized  in  brick  lined  drums.  It  was  then  passed  to  the 
iron  tube,  where  it  was  dead  roasted.  The  lime  was  sulphated  and  the 
iron  was  changed  to  the  ferric  state.  The  roasted  ore  from  the  revolv- 
ing drums  was  carried  by  a  chain  conveyor  directly  to  lead  lined  vats 
of  20-tons  capacity.  The  roasters  had  a  capacity  of  2  to  3  tons  each 
per  day. 

The  hot  ore  falling  from  the  conveyors  dropped  into  a  5  per  cent,  solu- 
tion of  sulphuric  acid.  The  solution  of  the  copper  took  place  rapidly. 
The  solution  from  the  leached  ore  contained  2.5  per  cent,  copper  and 
considerable  free  acid,  which  was  drawn  into  a  storage  vat.  The  tailings 
assayed  less  than  0.10  per  cent,  copper. 

The  solution,  as  drawn  from  the  ore  and  pumped  to  a  storage  tank, 
was  impregnated  with  sulphur  dioxide  gas,  made  by  burning  brimstone 
in  an  iron  pot  under  blast.  The  copper  solution  in  the  storage  tank, 
inipregnated  with  sulphur  dioxide,  was  then  flowed  into  electrolyzers, 
which  were  arranged  in  cascade  series  so  that  the  solution  could  flow 
from  one  to  the  other.  Sulphur  dioxide  was  also  blown  through  the 
electrolyte  in  the  electrolyzers  by  means  of  perforated  hard  rubber  tubes, 
which  in  addition  to  supplying  the  necessary  sulphur  dioxide,  agitated 
the  electrolyte,  and  thus  gave  the  desired  circulation. 

The  sulphur  dioxide  protected  the  lead  anodes  from  peroxidation. 
The  anodes  were  gradually  converted  into  lead  sulphate,  but  the  sul- 
phatization  was  much  slower  than  the  peroxidation  which  would  have 
occurred  without  the  introduction  of  the  sulphur  dioxide. 

Large  quantities  of  sulphuric  acid  were  regenerated  by  the  use  of 
sulphur  dioxide.  The  sulphur  dioxide  also  acted  as  a  depolarizer,  thus 
reducing  the  necessary  voltage,  and  consequently  the  amount  of  power, 
in  the  electrodeposition. 

The  copper  was  precipitated  at  the  Intercolonial  plant  at  1.5  volts, 
with  a  current  density  of  6  amperes  per  square  foot,  and  electrodes  about 
1   1/2  in.  apart.     The  current  eflaciency  was  about  90  per  cent.     The 

'  Electrochemical  Industry,  April,  1903. 


cathodes  were  greased  and  graphitized.  The  electrolysis  was  conducted 
until  the  copper  contents  was  reduced  from  2.5  to  1  per  cent.  The 
electrolyzed  solution,  regenerated  in  acid  by  the  secondary  anode  reac- 
tions, and  still  containing  1  per  cent,  copper,  was  returned  to  the  leaching 
vats,  and  the  cycle  continued  indefinitely. 

Keith  Process.— In  the  Keith  process  the  electrode  area  is  increased 
in  the  different  cells,  as  the  electrolyte  becomes  improvished  in  the 
metal  being  deposited.     The  electrodes  of  each  cell  are  in  multiple  in 

Fig.  57. — Keith   process.     Tanks  arranged  so  that  there  is  a,  gradual  reduction  of 
current  density,  as  the  electrolyte  becomes  impoverished  in  copper. 

the  cell,  but  in  series  in  their  relation  to  all  other  cells.     The  strength 
of  current  is  the  same  in  all  the  cells. 

It  is  evident  that  a  solvent,  strong  in  the  metal  being  deposited, 
entering  the  first  cell,  will  admit  of  a  greater  current  density  in  producing 
reguline  metal  on  the  cathodes,  than  will  the  weaker  electrolyte  entering 
the  succeeding  cells  in  the  series.  The  deposition  of  metal  in  each  cell 
of  the  series  impoverishes  the  electrolj^e  which  enters  the  succeeding 
cell,    and    therefore    the    current    density,    must    be    correspondingly 


less  in  order  to  insure  a  reguline  deposit  of  copper.  To  effect  this  Keith 
increases  the  number  of  electrodes,  or  the  surface  area  of  the  electrodes, 
in  a  progressive  order  in  each  succeeding  cell  from  the  first  to  the  last  of 
the  series  so  that  the  current  density  will  be  approximately  proportional 
to  the  strength  of  the  solution  in  the  metals  being  deposited.     Fig.  57. 

Keith  Process  at  Arlington,  New  Jersey/ — The  ore,  containing  the 
copper  as  chalcocite,  malachite,  azurite,  and  cuprite,  was  crushed  to 
30  mesh,  and  roasted  in  a  mechanical  furnace  with  a  hearth  200  ft.  long 
by  10  ft.  wide  and  fired  by  coal  in  seven  fireplaces  arranged  along  its  sides. 
The  capacity  of  the  furnace  was  about  125  tons  per  day. 

From  the  roasting  furnace  the  ore  was  conveyed  to  four  leaching  vats, 
30  ft.  in  diameter  and  6  ft.  deep,  holding  from  125  to  150  tons  of  the 
roasted  ore.  In  the  bottom  of  each  tank  there  was  a  filter  bed  of  coarsely 
crushed  rock,  covered  with  canvas,  which  was  caulked  tightly  around  the 
edges,  and  around  the  outlets  through  which  the  tailings  were  sluiced 
after  the  copper  was  extracted. 

From  the  center  of  the  bottom  of  each  tank,  under  the  filter  bed,  a 
pipe  with  a  stop  cock  served  through  which  to  draw  off  the  solvent  to  the 
electrolyte  cells.  The  first  solvent,  consisting  of  sulphuric  acid  and  largely 
of  ferric  sulphate,  took  up  nearly  all  the  copper  and  run  out  as  cupric 
sulphate,  and  decreased  in  amount  as  the  ore  became  improvished 
and  the  extraction  completed.  The  tanks  were  then  sluiced  out  and 

The  deposition  vats  were  rectangular  boxes  of  wood  with  a  suitable 
lining.  They  were  arranged  so  that  the  electrolyte  flowed  from  No.  1 
through  the  series  and  finally  out  of  No.  128,  depleted  of  its  copper,  into 
a  large  sump  tank,  from  which  is  was  pumped  back  into  the  stock  tank 
for  reuse. 

The  deposition  vats  were  so  set  at  different  elevations  above  the  floor 
building  that  the  electrolyte  run  by  gravity  from  one  to  the  next,  and  so 
on'through  the  series  to  the  end.  The  vats  were  arranged  in  six  rows, 
and  one  of  each  row  at  greater  elevation  than  the  end  next  to  it  of  the 
preceding  row,  it  was  necessary  to  raise  the  electrolyte  from  one  row  to 
the  next  at  those  points.     This  was  done  by  means  of  air  lift  pumps. 

The  electrical  connections  were  as  follows:  the  electrodes  in  each  vat 
were  connected  in  multiple,  and  the  several  vats  of  the  electrodes  con- 
nected in  a  series  of  128.  But  for  the  purpose  of  insuring  a  retrogressive 
decrease  of  current  density  at  the  electrodes  of  each  vat,  after  the  first  of 
the  series,  the  number  of  electrodes  was  progressively  increased,  from  the 
first  to  thp  last  of  the  series.  The  current  was  the  same  for  each  vat  of 
the  series,  but  the  current  density  was  less  and  less  in  decremental  order 
from  the  first  to  the  last  of  the  series,  to  compensate  for  the  decrease 

'  American  Inst.  Electrical  Eng.,  1902  Meeting,  S.  N.  Keith. 


of  copper  in  the  electrolyte  in  its  course  through  the  vats.  The  current 
density  was  from  15  to  20  amperes  per  square  foot  of  cathode  area  with 
a  6  per  cent,  copper  solution,  provided  proper  circulation  and  sufficiently 
rapid  movement  of  the  electrolyte  was  kept  up  between  the  anode  and 
cathode  surfaces. 

During  the  electrolysis  there  was  a  counter-electromotive  force  of  1.6 
volts  in  each  of  the  series.  The  generator  was  operated  at  approximately 
249  volts,  which  gives,  for  128  cells,  1.87  per  cell.  The  voltage,  per  cell, 
required  was  1.87—1.60  =  0.27  volt  for  resistance  of  conductors,  etc. 

Primarily  both  the  anodes  and  cathodes  were  of  sheet  lead,  but  under 
electrolytic  action  the  anodes  became  coated  with  PbOz  and  the  cathodes 
with  copper.  As  soon  as  the  copper  deposit  reached  a  thickness  of  card- 
board it  was  stripped  off  each  lead  cathode,  which  was  then  replaced  in 
its  cell,  and  the  two  copper  sheets  thus  produced  had  connections  rivited 
on  them,  and  were  then  rehung  as  cathodes  in  some  of  the  vats,  where 
they  remained  a  sufficient  length  of  time  to  receive  the  desired 

The  Siemens -Halske  Process. — In  this  process  the  ore  is  finely  ground 
and  roasted  at  a  moderate  temperature  in  such  a  way  that  the  iron  is 
almost  completely  oxidized,  while  the  copper  is  contained  in  the  roasted 
material,  partly  as  copper  sulphate,  and  partly  as  copper  oxide,  but 
principally  as  cuprous  sulphide,  CujS.  It  is  stated  by  the  inventors  that 
roasting  is  not  necessary  with  all  ores.  The  roasted  ore  is  then  treated  at 
a  temperature  of  about  90°  C.  (194° F.),  with  a  solution  of  ferric  sulphate, 
Fe2(S04)3,  to  which  is  added  a  little  sulphuric  acid. 

On  dissolving  the  copper,  the  ferric  sulphate  is  reduced  to  ferrous 
sulphate,  the  copper  going  into  solution  as  cupric  sulphate.  The  solution 
of  copper  and  ferrous  sulphate  is  then  led  into  the  cathode  compartment 
of  an  electrolytic  cell,  and  in  which  the  anodes  and  cathodes  are  separated 
by  a  permeable  diaphragm.  Part  of  the  copper  is  deposited  on  the  cath- 
ode, and  the  solution  then  circulates  through  the  diaphragm  to  the  anode. 
This  consists  of  carbon  rods.  At  the  anode  the  ferrous  sulphate  is  reoxi- 
dized  to  the  ferric  sulphate,  which  is  then  again  used  to  dissolve  copper 
from  new  charges  of  ore. 

The  chemical  reactions,  which  take  place  during  the  electrolysis,  and 
the  leaching  of  the  ore,  are  clearly  shown  by  the  following  equations: 

(1)  xH2SO,  +  2CuSO,  +  4FeSO,  =  2Cu  +  2Fe2(S04)3+xH2SO„ 

in  which  the  copper  is  electrolytically  precipitated,  and  the  ferrous 
sulphate  reconverted  to  the  ferric  sulphate,  at  the  anode.  The  electro- 
lyzed  and  regenerated  solution  is  then  returned  to  the  ore  and  the 
copper  dissolved; 

(2)  xH3SO,  +  Cu2S  +  2Fe2(SO,)3  =  2CuS04-F4FeSO,  +  S-FxH2S04. 


If  there  is  cupric  oxide  in  the  ore,  it  may  be  acted  upon  either  by  the 
sulphuric  acid  or  ferric  sulphate; 

(3)  CuO  +  H2S04  =  CuSO,  +  H20. 

(4)  3CuO+re2(S04)3  =  CuS04+FeA- 

A  comparison  of  equations  1  and  2  shows  that  if  the  copper  in  the 
ore  is  in  the  form  of  cuprous  sulphide,  the  electrolyte,  after  passing 
through  the  leaching  vats,  will  contain  exactly  the  same  quantity  of 
copper  sulphate,  ferrous  sulphate,  and  free  sulphuric  acid  as  it  did  prior 
to  electrolysis;  and  that  it  is,  therefore,  completely  regenerated,  and  may 
be  used  again  for  the  electrolytic  decomposition.  But  if  the  copper  is 
present  in  the  ore  partly  as  oxide,  it  is  evident  from  equations  3  and  4 
that  in  this  case  the  solution  will  be  richer  in  copper,  but  poorer  in 
respect  of  iron  and  sulphuric  acid  than  it  was  before  electrolysis.  These 
equations  do  not  take  into  account  the ,  possible  reactions  with  base 
elements  in  the  ore. 

If  a  solution  containing  cupric  and  ferrous  sulphates  is  electrolyzed 
in  the  presence  of  sulphuric  acid,  copper  is  deposited  in  preference  to  the 
iron.  If  the  electrolysis  is  performed  without  a  diaphragm  between  the 
electrodes,  the  ferrous  sulphate  is  oxidized  at  the  anode  to  ferric  sulphate, 
and  reduced  again  at  the  cathode  to  ferrous  sulphate.  This  represents 
a  waste  of  energy,  which  appears  as  heat,  and  the  electrolyte,  as  a  solvent 
for  copper,  is  not  much  improved.  Ferric  sulphate  is  a  solvent  of  copper 
from  its  oxide  and  sulphide  combinations;  ferrous  sulphate  is  not;  it  is 
therefore  desirable  to  have  as  much  as  possible  of  the  iron  in  the  solution 
in  the  ferric  condition,  before  again  applying  it  to  the  ore,  and  this  is 
brought  about  by  interposing  diaphragms  between  the  electrodes,  and 
then  passing  the  solution  from  the  cathode  to  the  anode  compartment, 
or  by  permeating  the  solution  through  the  diaphragm  from  the  cathode 
to  the  anode  compartment. 

The  scheme  in  the  Siemens-Halske  process,  at  first  contemplated 
roasting  the  ore  containing  the  copper  sulphides,  at  a  low  temperature,  so  as 
to  oxidize  the  sulphide  of  iron  which  it  contains,  to  ferric  oxide,  and  to  free 
the  cuprous  sulphide  originally  forming  a  constituent  of  copper  pyrites 
in  the  ore.  In  the  course  of  the  roasting  some  of  the  cuprous  sulphide 
is  converted  into  sulphate,  but  this  only  advances  the  process  by  the 
amount  of  copper  sulphatized. 

Copper-iron  sulphide,  as  occurring  in  nature  in  the  ore,  is  not  readily 
soluble  in  ferric  sulphate  so  that  an  impracticable  long  time  is  required 
to  effect  the  solution  of  the  copper,  even  with  exceedingly  fine  grinding, 
so  that  a  commercial  application  of  this  method  of  leaching  sulphide  ores 
is  not  feasible  under  existing  conditions.  It  is  impossible  to  perform 
the  delicate  roast  necessary  to  release  the  copper  sulphide  and  oxidize 
the  iron  sulphide. 



Free  copper  sulphides  and  oxides  react  with  ferric  sulphate  easily  and 
quicldy.  The  presence  of  a  large  quantity  of  ferrous  sulphate  in  the 
ferric  sulphate  solution  impairs  the  solution  of  the  copper  from  cuprous 
sulphide  by  means  of  ferric  sulphate. 

The  method  suggested  in  the  Siemens-Halske  process  for  roasting 
sulphide  ore,  so  that  the  main  quantity  of  the  iron  is  transformed  to 
oxide  while  the  larger  portion  of  the  copper  remains  as  cuprous  sulphide, 
is  quite  impossible.  Neither  can  satisfactory  results  be  obtained  by 
dead-roasting,  for  the  reason  that  at  the  temperature  required,  basic 
silicates  are  formed  by  means  of  a  combination  of  the  copper  oxide  with 
the  silicates  of  the  gangue,  and  perhaps  also  salts  of  the  type  FegO^  are 
formed  by  combination  of  the  oxides  of  copper  and  iron;  such  salts  are 
acted  upon  very  slowly  by  ferric  sulphate.  Cuprous  oxide  is  also  formed 
by  roasting  at  fairly  high  temperatures,  and  the  cuprous  oxide  so  formed 
is  not  readily  soluble  in  a  solution  of  ferric  sulphate. 

There  is  no  difficulty  in  making  the  copper  soluble  by  roasting  at  a 
low  temperature.  This  temperature  is  about  450  to  480°  C.  The 
copper  in  the  ore,  at  these  temperatures,  is  largely  converted  into  sulphate 
and  some  into  oxide.  There  would  appear  to  be  no  reason  why  the 
roasting  should  be  attempted  to  be  carried  on  as  originally  proposed  by 
the  inventors,  when  it  is  more  satisfactory  and  just  as  cheap  to  give  the 
ore  a  thorough  roast  at  a  low  heat. 

X  ' 

Fig.  58. — Horizontal  diaphragm  cell.     Used  in  the  Siemens-Halske  process. 

Various  difficulties  were  encountered  in  the  practical  operation  of  the 
Siemens-Halske  process,  principally  among  these,  was  the  indifferent 
nature  of  the  solvent  and  the  inability  to  obtain  suitable  anodes  and 
diaphragms.  The  anode  difficulty  has  not  yet  been  overcome,  as  this 
process,  in  common  with  all  other  sulphate  processes  is  still  laboring 
under  the  disadvantage  of  not  being  able  to  find  a  suitable  insoluble 
anode.  In  the  Siemens-Halske  process,  as  experimentally  carried  out 
some  years  ago,  carbon  was  used  as  the  anode,  but  it  was  not  at  all 

In  the  later  form  of  apparatus  used  in  the  electrolysis,  the  diaphragms 
and  electrodes  were  placed  horizontally,  and  constructed  as  shown  in 
Fig.  58.     The  electrolyzer,  E,  is  divided  horizontally  into  two  compart- 




ments  by  an  asbestos  diaphragm,  D.  In  the  upper  compartment  is  the 
cathode  C,  and  in  the  lower  compartment  the  anode  A.  The  cathode  may 
be  made  of  a  thin  sheet  of  copper  and  the  anode  of  carbon  or  of  sheet 
lead.  The  solution  from  the  ore  is  introduced  at  K,  and  drawn  off  at  Y, 
the  rate  of  flow  being  adjusted,  so  that  it  passes  slowly  and  continuously 
through  the  permeable  diaphragm  D,  and  is  in  contact  with  the  electrodes 
successively  for  a  sufficient  time  to  allow  the  deposition  of  most  of  the 
copper  in  the  upper  compartment,  and  of  the  oxidation  of  the  ferrous 
sulphate  to  ferric  sulphate  in  the  lower  compartment.  The  electrolyzed 
solution  as  withdrawn  from  the  anode  compartment  is  returned  to  the 

General  arrangement  of  a  Siemens-Halske  plant. 

Fig.  69  shows  a  complete  outline  plant  for  the  Siemens-Halske  pro- 
cess. A  is  the  storage  tank  for  the  solution  to  be  electrolyzed.  The 
solution  passes  through  the  pipe  B  into  the  bath  C,  flows  first  into  the 
cathode  division  k,  and  then  through  the  filter  into  the  anode  division  a, 
from  which  the  escape  pipe  D  leads  it  to  the  pipe  G,  which  conducts  it  into 
the  solution  tank  H.  Here  it  comes  in  contact  with  the  ore  to  be  leached, 
which  has  been  ground  in  the  ball  mill  E.  After  the  copper  has  been 
dissolved,  the  mixture  of  exhausted  ore  and  liquor  runs  into  the  vacuum 
filter  K.  The  solution  aspirated  through  the  filter  is  again  conveyed 
to  the  storage  tank  A  by  the  pipe  M. 

The  theoretical  voltage  required  in  the  Siemens-Halske  process,  to 
precipitate  the  copper  and  convert  the  corresponding  amount  of  ferrous 
sulphate  to  ferric  sulphate  is  0.36  volt.     The  inventors  considered  that 


0.7  volt  would  give  a  current  density  of  the  required  strength  for  practical 
operations.  In  the  experimental  tests  it  varied  from  0.75  to  1.8  volts. 
The  process  is  nowhere  now  in  practical  use. 

M.  DeKay  Thompson  Jr.'s  Experiments  on  the  Siemens -Halske 


Interesting  experiments  on  the  Siemens-Halske  process  were  made 
by  M.  DeKay  Thompson  Jr.'  to  determine  the  solvent  action  of  ferric 
sulphate  on  copper  compounds  likely  to  occur  in  raw  or  roasted  ore,  and 
also  to  determine  the  efficiency  of  the  electrolytic  precipitation.  His  con- 
clusions may  be  summarized  as  follows: 

Cupric  Oxide,  CuO. — 1.  The  investigation  of  the  action  of  ferric  sul- 
phate on  cupric  oxide  leads  to  the  conclusion  that  the  reaction  between 
the  two  is  probably  represented  by  the  equation: 

3CuO  -l-FejCSO,)  3 =3CuS04 +Fe203, 

when  the  two  are  present  in  equivalent  quantities.  When  this  is  not  the 
case,  basic  salts  are  formed  to  a  considerable  extent. 

2.  Copper  sulphate  is  precipitated  by  copper  oxide.  When  the 
amount  of  oxide  is  equivalent  to  the  amount  of  copper  in  the  solution, 
the  precipitation  is  only  partial;  in  the  presence  of  a  large  excess  of  copper 
oxide  it  is  complete. 

3.  A  metathesis  takes  place  between  copper  oxide  and  ferrous  sul- 
phate, analogous  to  that  with  ferric  sulphate.  Ferrous  oxide  and  copper 
sulphate  are  the  resulting  products. 

4.  Under  certain  conditions  all  iron  and  copper  salts  could  be  thrown 
out  of  solution. 

Cuprous  Oxide,  CujO. — 1.  The  results  of  the  experiments  with  cuprous 
oxide  are: 

1.  Cuprous  oxide  reduces  ferric  sulphate  to  ferrous  sulphate  according 
to  the  equation, 

CujO  +Fe2(S0,)  3  +  H^SO,  =  2CuS0, +2FeS0,  +  H^O, 

and  both  ferrous  and  ferric  iron  is  precipitated. 

2.  Cupric  sulphate  does  not  act  on  cuprous  oxide. 

3.  Ferrous  sulphate  does  not  act  on  cuprous  oxide. 

Cuprous  Sulphide,  CujS. — The  results  of  the  experiments  with  cuprous 
sulphide  showed: 

1.  The  verification  of  the  equation, 

Cu,S  -F  2Fe2  (SO  J  3  =  2CuS0 ,  +  4FeS0 ,  +  S. 

2.  The  FeSO^  formed  has  no  effect  on  cuprous  sulphide. 

3.  The  CUSO4  formed  has  no  effect  on  cuprous  sulphide. 

'  Electrochemical  Industry,  June,  1904. 


4.  Sulphuric  acid  dissolves  cuprous  sulphide  in  the  presence  of 
oxygen  very  slowly.  The  presence  or  absence  of  sulphuric  acid  in  the 
solvent  is  therefore  of  very  little  consequence  as  far  as  the  CujS  is 

Cupric  Sulphide,  CuS.— The  following  seems  to  be  the  simplest  reac- 
tion that  can  take  place  between  copper  sulphide  and  ferric  sulphate: 

CuS+Fe,(SOj3  =  CuSO,  +  2FeSO,  +  S. 

In  the  tests  some  sulphur  was  set  free,  but  the  amount  of  copper  dis- 
solved was  in  excess  of  that  called  for  by  the  above  equation.  This 
excess  was  evidently  due  to  the  oxidation  of  the  copper  sulphide  to 
sulphate.  To  show  this,  tests  were  then  made  with  sulphuric  acid  of 
1.2  sp.  gr.,  with  a  1.15  per  cent,  solution  of  cupric  sulphate,  and  with  a 
0.55  per  cent,  solution  of  ferrous  sulphate.  In  all  cases  about  the  same 
amount  of  copper  was  dissolved,  which  shows  that  it  was  due  to  the 
oxygen  present.  Copper  sulphide  is  therefore  not  dissolved  by  any  of 
these  reagents. 

Many  copper  ores  contain  iron,  which  is  changed  over  to  oxide  on 
roasting.  The  reaction  that  takes  place  between  ferric  oxide  and 
ferric  sulphate  is  to  precipitate  iron  from  the  solution  as  a  basic  salt. 

Experiments  were  then  made  with  a  copper  ore  having  the  following 

Cu,  29 .  99  per  cent. 

Fe,  27 .  89  per  cent. 

S.  33 .  32  per  cent. 

SiOj,  9.75  per  cent. 

100.95  per  cent. 

The  composition  of  the  mineral  in  this  ore  corresponds  closely  to  that  of 
copper  pyrites. 

The  first  experiment  showed  that  the  ore  was  scarcely  attacked  by 
a  5  per  cent,  solution  of  ferric  sulphate.  The  powdered  ore,  crushed  to 
80  mesh,  was  then  roasted  for  several  hours.  After  roasting  it  was  found 
to  contain  28.8  per  cent,  copper,  and  18.9  per  cent,  sulphur,  and  the  iron 
was  computed  to  be  27.0  per  cent.  Some  of  the  iron  and  copper  was 
oxidized,  so  that  7.6  per  cent,  of  the  ore  was  copper  sulphate,  and  2.3 
per  cent,  iron  sulphate.  Solution  tests  showed  that  not  much  more  than 
75  per  cent,  of  the  copper  in  this  roasted  ore  could  be  dissolved  from  this 
sample  with  a  ferric  sulphate  solution  containing  2.833  grm.  of  iron  in 
50  c.c.  of  solution. 

A  second  set  of  experiments  were  made  with  a  smaller  amount  of  the 
ore,  and  the  same  volume  of  solvent.  In  this  case  a  little  more  copper 
was  dissolved.  These  results  confirmed  those  previously  obtained. 
Experiments  were  then  made  to  see  if  more  of  the  copper  could  be  dis- 


solved  out  of  the  ore  already  treated  by  using  fresh  portions  of  the  solvent. 
After  from  three  to  five  treatments  in  this  way,  the  ultimate  extraction 
was  81.0  per  cent.  The  residues  from  these  experiments  were  collected 
and  roasted  again.  It  was  found  that  by  this  means  91.8  per  cent,  of 
the  copper  still  remaining  in  the  ore  was  dissolved  by  the  solvent  in  one 

For  this  reason  the  ore  used  in  the  above  experiments  was  roasted 
again.  The  copper  then  amounted  to  30.3  per  cent.  This  ore,  when 
agitated  for  1  hour  with  a  6  per  cent,  solution  of  ferric  sulphate  showed 
an  extraction  of  93.9  per  cent,  of  the  copper.  In  5  hours,  under  similar 
conditions,  98.9  per  cent,  of  the  copper  was  dissolved. 

These  experiments  show  that  copper  can  be  dissolved  by  ferric 
sulphate  pretty  completely  as  soon  as  the  ore  is  sufficiently  roasted. 
The  question  is  what  are  the  compounds  formed  by  roasting,  which  are 
so  much  more  soluble  than  the  unroasted  ore.  It  seems  that  this  is  due 
to  the  formation  of  copper  oxide.  If  this  is  the  case,  just  as  much  copper 
would  be  dissolved  by  sulphuric  acid.  To  test  this  conclusion  some  of 
the  ore  was  similarly  treated  with  sulphuric  acid  of  1.2  sp.  gr.  for  5  hours. 
This  showed  an  extraction  of  97.7  per  cent,  of  the  copper.  It  was  found, 
however,  in  the  study  of  cuprous  sulphide,  that  this  was  easily  extracted 
by  ferric  sulphate,  and  this  is  exactly  what  was  under  investigation  in 
these  experiments,  combined  with  sulphide  of  iron.  It  must  be  this 
double  combination,  therefore,  that  prevents  the  copper  from  being  ex- 
tracted by  the  ferric  sulphate. 

The  results  of  the  experiments  with  the  ore  may  be  summarized  as 
follows : 

1.  Copper  pyrites  is  not  appreciably  attacked  by  ferric  sulphate. 

2.  Roasting  so  changes  the  ore  that  nearly  all  the  copper  can  be 
extracted  by  either  sulphuric  acid  or  by  ferric  sulphate.  This  makes  it 
seem  probable  that  the  roasting  changes  the  copper  largely  over  to  oxide. 

The  Electrolysis. — The  electrolysis  comprises  two  reactions,  the  reduc- 
tion of  the  copper  at  the  cathode  and  the  oxidation  of  the  iron  at  the 
anode.  These  were  investigated  separately.  The  vessel  used  was 
a  copper  voltameter  jar  divided  into  three  narrow  compartments  by  two 
clay  diaphragms  made  fast  to  the  glass  sides  with  paraffine.  Theanode 
and  cathode  were  in  the  outer  compartment,  while  the  inner  one  pre- 
vented diffusion  from  one  electrode  to  the  other. 

In  the  experiments  on  the  deposition  of  copper  a  lead  plate  was  used 
as  anode,  and  the  anode  compartment  was  filled  with  sulphuric  acid. 
The  cathode  compartment  was  filled  with  a  solution  of  the  following 

5  per  cent,  ferrous  iron. 

3.5  per  cent,  copper. 

2.5  to  3.0  per  cent,  sulphuric  acid 



The  same  solution  was  contained  in  the  middle  compartment.  During 
electrolysis  the  liquid  in  the  cathode  compartment  was  stirred  by  a  cur- 
rent of  carbon  dioxide. 

The  object  in  these  experiments  was  to  determine  how  poor  the  solu- 
tion may  become  in  copper  without  affecting  the  character  of  the  copper 
deposited  at  a  given  current  density.  This  was  to  be  determined  for 
the  different  current  densities.  For  this  purpose  the  gain  in  weight  of 
the  copper  cathode  was  compared  with  the  copper  deposited  in  a  copper 
voltameter.  The  accompanying  table  gives  the  results  obtained  with  a 
current  density  of  0.98  ampere  per  square  decimeter.  (9.1  amperes  per 
square  foot.) 

Time  in 

Cu  deposited  in  voltameter 
for  1/2  hour 

Cu  deposited  in  cell  for  1/2  hour 

Per  cent,  yield 









1    1/2 








2  1/2 








3  1/2 








4  1/2 



After  4  hours  the  copper  commenced  to  be  spongy.  The  cathode  solution 
was  then  analyzed  and  was  found  to  contain  0.72  per  cent,  copper.  In 
another  similar  experiment  with  the  same  current  density,  the  copper 
did  not  begin  to  be  spongy  till  the  strength  of  the  solution  had  reached 
0.38  per  cent,  copper. 

In  still  another  experiment  with  a  current  density  of  1.8  amperes  per 
square  decimeter  (16.8  amperes  per  square  foot)  the  copper  began  to  get 
spongy  at  the  end  of  the  second  hour.  The  electrolysis  was  therefore 
discontinued  and  the  solution  in  the  cathode  compartment  analyzed. 
It  contained  0.98  per  cent,  copper.  The  current  density  was  then  reduced 
to  0.47  ampere  per  square  decimeter  (4.4  amperes  per  square  foot)  and 
the  electrolysis  continued.  The  copper  came  down  in  good  form  till  the 
concentration  of  copper  was  0.05  per  cent,  in  the  solution.  Two  other 
experiments  were  carried  out  with  current  densities  of  3.4  amperes  and 
2.6  amperes  per  square  decimeter,  respectively.  (31.6  and  24.2  amperes 
per  square  foot.)  In  the  first  case  the  copper  was  deposited  in  a  spongy 
form  immediately.  The  same  was  true  in  the  second  case  though  not  in 
so  marked  a  manner.  Some  of  the  spongy  copper  was  tested  for  iron, 
but  this  impurity  was  not  present. 

Oxidation  of  the  Iron  at  the  Anode. — In  all  of  the  experiments  to 
determine  the  oxidation  of  the  iron  at  the  anode,  the  yield  for  the  first 


few  hours  amounted  to  more  than  100  per  cent.  It  seems  probable,  there- 
fore, that  if  the  disturbing  causes  were  removed,  the  yield  would  not  fall 
much  below  100  per  cent.  At  the  end  of  the  first  period  the  solution 
contained  2.1  per  cent,  ferrous  sulphate.  Therefore  the  conclusion  may 
be  drawn  that  starting  with  a  solution  containing  5  per  cent,  ferrous  iron, 
this  may  be  oxidized  with  approximately  100  per  cent,  yield,  till  the 
solution  contains  only  2  per  cent,  of  ferrous  iron,  using  a  current  density 
of  from  0.3  to  0.5  ampere  per  square  decimeter  (2.79  to  4.64  amperes  per 
square  foot) .  When  the  solution  contains  between  2  per  cent,  and  1  per 
cent,  oxidizable  iron,  the  yield  would  still  be  about  90  per  cent.  When 
the  concentration  is  still  more  diminished,  the  evolution  of  gas  becomes 
strong  and  the  yield  falls  off  correspondingly. 

Siemens-Halske  Process  in  Spain. — The  Siemens-Halske  process  was 
installed  for  practical  operation  in  Southern  Tyrol  in  Spain,  but  the  results 
were  unsatisfactory,  owing  mostly  to  the  indifferent  nature  of  the  solvent. 
The  ore  contained  the  copper  in  form  of  the  compound  CujS.FeS.FeSj, 
which  was  found  difficult  to  dissolve,  either  raw  or  roasted,  to  encourage 
further  operations.  Later,  the  ore  was  roasted  at  a  very  low  tempera- 
ture, and  leached  with  sulphuric  acid,  and  the  copper  in  the  solution 
crystallized  to  copper  sulphate,  for  which  there  was  a  good  local  market. 

Experiments  at  the  Ray  Mines,  Arizona. — W.  L.  Austin  gives  an  ac- 
count of  experiments  made  at  the  Ray  Mines,  by  W.  Y.  Westervelt  in  the 
summer  of  1905,  which  are  interesting.' 

"Westervelt,  after  making  preliminary  wet  concentration  tests  which  gave 
unsatisfactory  results,  instituted  leaching  experiments  on  the  Ray  ore.  These 
were  sufficiently  encouraging  to  warrant  the  construction  of  a  small  plant  on  the 
property  for  the  purpose  of  carrying  out  further  investigations. 

"  The  ore  from  the  Ray  mines  consists  of  disseminated  sulphides  in  a  porphyry 
—or  schistose — gangue,  and  is  thought  to  average  about  2.22  per  cent,  copper, 
accompanied  by  little  or  no  precious  metals.  Immense  quantities  of  this  ore 
are  known  to  exist. 

"At  Ray,  leaching  with  sulphuric  acid  was  found  to  remove  a  small  percent- 
age of  the  copper  from  the  crude  ore  but  did  not  attack  that  which  was  shut  up 
in  the  sulphides.  However,  practically  all  of  the  copper  present  could  be  readily 
brought  into  solution  by  treatment  with  a  hot-acid-solution  of  ferric  sulphate, 
and  investigations  along  t