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P E B P A C E. 

The lectures on "Modern Copper Smelting" embodied in this 
volume were delivered at the University of Birmingham to the 
Senior Students in the School of Metallurgy and to others 
interested in the subject. 

They are based largely upon the results of a study of the 
practice as conducted at a number of the best organised smelters 
and refineries in the United States of America, at which the author 
has had the opportunity of spending some considerable time, and 
it has been felt that there exists a scope, particularly on this side 
of the Atlantic, for a compact volume dealing broadly with the 
principles underlying Modern Copper Smelting, illustrated with 
such examples of working practice from personal observation. The 
subject-matter of the Lectures has been extended by the addition 
of an Introduction on the History, Uses, and General Metallurgy of 
Copper as applied to Modern Practice. 

The Copper Industry is already fortunate in the literature at 
its disposal. It possesses standard works of reference through the 
publication of Dr. Peters' classical volumes on the Principles of 
Copper Smelting, and more recently (during the preparation of the 
present work) of the volume on the Practice of Copper Soneltirig — 
works which have done much to raise copper smelting to a science. 
The industry is being rendered invaluable service by the Technical 
Societies and Technical Press, whose publications furnish an admir- 
able record of the constant advance in the theory and practice of 
the art. Use has been made of these sources of information in the 
present work, and lists of such references are appended to each of 
the Lectures. 

Grateful acknowledgment is made to several authors and editors 
who have given permission for the reproduction of illustrations or 
for the inclusion of references : — Dr. Peters, Professor Gowland, Mr. 


Hughes, the Editors of the Engineering and Mining Journalr 
Mineral Industry, Mines and Minerals, and others. The Institution 
of Mining and Metallurgy, Messrs. Chambers Bros., The Traylor 
Engineering Co., and the Power and Mining Machiner}^ Co. have 
very kindly provided blocks for several of the illustrations; the 
Anaconda Copper Mining Co. furnished a set of photographs, whilst 
Figs. 8, 37, and 76 have been reproduced by permission of the 
American Institution of Mining Engineers. 

To the Superintendents and Staffs of the several smelters where 
opportunities were so freely given for studying modern practice, and 
particularly to Mr. E. P. Mathewson at Anaconda, Montana, to Mr. 
J. Parke Channing at the Tennessee Copper Company's Smelter, and 
to Mr. W. H. Freeland at Ducktown, Tennessee, the author desires to 
express his appreciation for much valued information and many other 
kind services. The frequent references made in this book to the 
organisation and the methods employed at these works is not only 
a tribute to the useful information freely imparted, but is also due to 
the fact that such features are so thoroughly representative of the 
most advanced practice in copper smelting upon a large scale and of 
the direction in which all modern work is undoubtedly tending. 

The author further thanks Professor Turner of Birmingham 
University for his interest in this volume, Mr. Frank Levy for 
reading the proofs, and the publishers, Messrs. Charles Griffin & Co., 
Ltd., for the care taken in the preparation and production of the 

University of Birmingham, 
May, 1912. 



History of Copper — Development of the Copper Industry — Progress of 
Smelting Practice— Price and Cost of Production of Copper— Copper 
Statistics, . . . . . . . . . 1-17 


The Uses of Copper : as Metal and as Alloy — The Physical Properties of 
Copper — Efifects of Impurities — Mechanical Properties — Chemical 
Properties, ... . . . . . 18-34 


Compounds of Copper— Copper Mattes — The Varieties of Commercial 

Copper — Ores of Copper — Preliminary Treatment of Ores — Sampling, 35-50 


Modern Copper Smelting Practice — Preliminary Treatment of Ores : 
Concentration, Briquetting, Sintering — The Principles of Copper 
Smelting — Roasting, ....... 51-80 


Keverberatory Smelting Practice : — Functions of the Reverberatory Furnace 
— Requirements for Successful Working — Principles of Modern Rever- 
beratory Practice — Operation of Modern Large Furnaces — Fuels for 
Reverberatory Work ; Oil Fuel ; Analysis of Costs— Condition of the 
Charge, 81-112 




Blast-Purnace Practice :— Functions of the Furnace— Reduction Smelting— 
Oxidation in the Furnace— The Pyritic Principle— Features of Modern 
Working: Water-Jacketing, Increase in Furnace Size, External 
Settling— Constructional Details of the Furnace, . . . 113-145 


Modern Blast-Furnace Practice (continued) :— Charge Calculations — Working 
— Disposal of Products — Pyritic Smelting — Sulphuric Acid Manu- 
facture from Smelter Gases, ...... 146-191 


The Bessemerising of Copper Mattes :— Development of the Process — The 
Converter — Converter Linings — Grade of Matte — Operation of the 
Process — Systems of Working, ...... 192-216 


The Purification and Refining of Crude Copper : — Preliminary Refining and 
Casting into Anodes — Electrolytic Refining — Bringing to Pitch, and 
Casting of Merchant Copper, ...... 217-243 

Index, .......... 245-259 


Frontispikce — The Colour of the Converter Flame during the Bessemerising of 

Copper Matte. 


Fig. 1. — Fluctuations in the Price of Best Select Copper, . . . .12 

-Annual Production of Copper, . . . . . .16 

-Equilibrium Diagram, Cu-Zn Series, . . . . .22: 

-Influence of Arsenic and Antimony on the Electrical Conductivity of 

Copper, ......... 25 

-Relations of Copper and Oxygen, ...... 27 

-Microstructure of Copper containing Oxygen (Heyn), . Plate to face 28- 
-Relations of Copper and Arsenic, ...... 2^ 

-Freezing-Point Curve of Iron-Copper Sulphides (Mattes), . . 38- 

-Outline of Sampling Scheme, Anaconda, ..... 48 

-Section through Sampling Mill, ...... 48- 

-Brunton Sampler, ........ 49" 

-Outline of Smelting Scheme at the Anaconda Smelter, Montana, U.S.A., 54 
-Sketch Plan of Briquetting Plant, ...... 56 

-Section through Auger-Former, showing Briquetting Mechanism of 

Chambers' Machine, ....... 56 

-Chambers' Briquette-making Machine, . . . Plate to face 58 

-Dwight-Lloyd Sintering Machine, ...... 60 

-O'Harra Furnace (Fraser-Chalmers), illustrating Principle of Mechanical 

Rabbling by Travelling Ploughs, ..... 70 

,, 18. — Section through Mechanically Rabbled Roaster Furnace (illustrating 

Improvements for Protecting Driving Mechanism), . . .71 

,, 19.— MacDougal Roaster — Vertical Section, . . . . .74 

,, 20. — Herreshof Furnace— Section indicating Connections for Cooling Rabbles 

and Spindles, ........ 74 

,, 21. — Spindle Connections and Guide Shields of Evans-Klepetko Roasters, . 76 
„ 22.— Rabble-blades and Bases, ....... 77 

,,23. — Development of the Reverberatory Furnace (Gowland), . . .90 

,, 24. — iJraft Pressure Record of Anaconda Reverberatory Furnace, . . 94 

,, 25.— Skimming Reverberatory Furnace, Anaconda, . . Plate to face 96 

,, 26. — Transverse Section of Modem Reverberatory Furnace, Anaconda, 

indicating Foundations, Hearth, and Bracing, . . .96 

,,27. — Reverberatory Furnace under Construction, . . Plate to face 96 



















Fig. 28.— Sectional Plan and Elevation of Reverberatory Furnace at Anaconda, 98 

,, 29.— Fire-box End of Reverberatory Furnace, showing massive Bracing, 

Charge Bins, and Charging Levers, Anaconda, . Plate to face 100 

,, 30.— Interior of Reverberatory Furnace (looking towards Skimming Door), 

showing Expansion Spaces in Roof, and Charging Holes, 

Anaconda, ...... Plate to face 100 

,, 31.— Shelby Oil-Burner for Reverberatory Furnace Use, . . . 106 

,, 32. — Modern Blast- Furnace Shell of Sectioned Jackets (P. & M. M. Co.), 

Plate to face 122 
,, 33. — Blast Furnaces under Construction, showing Fixing of Jackets, Bottom 

Plate, Method of Support, Sectioning, etc. (T. E. Co.), Plate to face 124 
^, 34. — Development of the Blast Furnace (Gowland), .... 126 
,, 35. — Plan of 51-foot Blast Furnace, Anaconda, indicating Position of 

Crucibles, Spouts, and Connecting Bridge between Old Furnaces, . 128 
^, 36. — Longitudinal Section and Part Elevation of S7-foot Blast Furnace, 

Anaconda, indicating Crucibles of Old Furnaces, Bridge, and 

Jacketing, ........ 128 

,, 37. — Copper Contents in the Slags accompanying Mattes of Various Grades, 132 
,, 38. — Water- Jacketed Blast Furnace, lower portion indicating Air and 

Water Connections, Bottom Supports, End Slag Spouts, etc. 

(P. & M. M. Co.), ..... Plate to face 134 
,, 39. — Tapping Breast of Blast Furnace, Cananea, . . . .136 

^, 40. — Rivetted Steel Water- Jacket, showing Tuyere Holes and Water 

Inlets, etc. (P. & M. M. Co. ), . . . . . 137 

,, 41. — Transverse Section through Modern Blast Furnace, showing Arrange- 
ments of Boshed Lower Jackets, Upper Jackets, and Plates, Stays 

and Supports, etc., ....... 138 

,, 42. — Interior of Anaconda Blast Furnace, showing Jacketing, Tuyere 

Holes, and Bridge, ..... Plate to face 138 
^, 43. — Showing Upper Jackets, Apron and Mantle Plates and Superstructure 

of Blast Furnace, Anaconda, .... Plate to face 140 
-,, 44. — Charging Blast Furnaces, Anaconda, . . . Plate to face 140 

^, 45.— Blast-Furnace Shell, with Air Connections (P. & M. M. Co.), . . 142 

^, 46. — Details of Tuyere, Cananea Blast Furnace, . . , . 142 

-,, 47. — V-Shaped Charging Car, indicating Mechanism for Release and 

Tilting, ........ 153 

^, 48.— End View of Blast Furnace, showing Tilting of Charge Car, Anaconda, 155 
,, 49. — Hodge's Charging Car, . . . , . . . 155 

^, 50.— Freeland Charging Machine (D. S. C. & I. Co.), . . .157 

„ 51. — Freeland Charger — Details, . . . . . .157 

.,, 52 —Slag Spout, showing Method of Trapping Blast, also Replaceable 

Nose-Piece of Spout (A), . . . . . . 159 

.,, 53. —Details of Slag Spout, Cananea, ...... 161 

yy 54.— Slag Spout, showing Method of Support, . . . . 161 

,, 55.— General View of Settler (T. E. Co.), ..... 163 

„ 56.— Method of Lining Settler, Cananea, ..... 163 



Fig. 57. — Arrangement for Matte and Slag Discharge from Settlers (T. C. C), 
,, 58. — Tap-hole Casting and Detail for Settlers, 

., 59. — Anaconda Blast Furnace (51 feet long), showing Settlers, Plate to face 
., 60.— Hoppers of Flue-Dust Chambers and Tracks for Cars underneath, 
,, 61.— Slotted Tuyeres, 12 inches by 4 inches (T. C. C), 
,, 62. — Sectional Elevation and Plan of Barrel-Shaped Silica-Lined Con 

verter (Peters), ...... 

., 63. — Latest Form of Silica- Lined Barrel Converter, . 

,, 64. — Longitudinal Section of Basic-Lined Converter, 

,, 65. — Basic-Lined Converter, indicating Tuyeres, Lining, &c., 

„ 66.— Composition of a Charge during Bessemerising Operation, 

,, 67. — Pouring Slag, Anaconda, ...... 

,, 68. — General View of Converter Shop, Anaconda, . . Plate to face 

,, 69. — Sectional Plan, Elevation, and Transverse Sections of Refining and 

Anode-Casting Furnace, Anaconda (Peters), . . . . 

,, 70. — Indicating Tilting and Pouring Mechanism of Ladle of Casting and 

Refining Furnaces, ....... 

,, 71. — Walker's Anode-Casting Machine, . . . Plate to face 

,, 72. — General View of Tank-room of Electrolytic Refinery (Perth Amboy, 

N.J.), ....... Plate to face 

,, 73. — Indicating Methods of Suspending and Connecting Electrodes (Perth 

Amboy, N.J. ), . ....... 

„ 74. — Indicating Connections for Circulation of Electrolyte (Barnett), 

„ 75. — Tank-house, showing Anode Crane (Ulke), . . . . 

„ 76. — Microstructure of Commercial Copper containing Oxygen (Hofman), 

Plate to face 














I. The Production of Copper, .... 

II. North American Production of Copper, 

lit. Influence of Impurities on the Electrical Conductivity of 

IV. Analysis of Various Commercial Coppers, . 

V. Development in Size of the Reverberatory Furnace, 

VI. Daily Reports. Reverberatory Furnaces, . 

VII. Daily Assay Report. Reverberatory Furnaces, . 

VIII. Monthly Report. Reverberatory Furnaces, 

IX. Effect on Coke Consumption of Increased Sulphur in the 
Charge, ..... 

X. Blast- Furnace Charge Calculations, . - 

XI. Typical Charging Tables at Pyritic Smelter, 

XII. Changes in Composition during Bessemerising, 














History of Copper— Development of the Copper Industry 
— Progress of Smelting" Practice— Price and Cost of 
Production of Copper— Copper Statistics. 

The History of Copper, — Copper was probably the earhest metal 
commonly employed by mankind. It occurs in the native condition 
in various parts of the world, and the natural product thus required 
no metallurgical treatment prior to use. Its malleability and the 
property of being readily toughened by simple mechanical treatment 
were also factors which account for the discovery of its general use- 
fulness in such primitive times. 

Although silver and gold were possibly known even earlier, these 
metals appear to have been employed chiefly for ornamental purposes, 
and as tokens, rather than for general service. 

The alloy of copper and tin, known as bronze, was the first metallic 
combination in common use by man ; its employment was so charac- 
teristic in prehistoric times, that archaeologists assign to one of the 
epochs the name of the Bronze Age. As is well known, archaeological 
time is marked by a series of ages, in which the use, first of stone, 
then of bronze, and ultimately of iron for the manufacture of tools 
and implements, indicate the development of industrial culture. The 
dates which can be assigned to those periods vary with the locality ; 
the races in the more Northerly latitudes being later in their develop- 
ment. In our own country, the Stone Ages may be said to date 
from 3000 B.C. down to 1000 B.C., and the Early and Late Bronze 
Ages from 1000 B.C. to 500 B.C., and from 500 B.C. to the commence- 
ment of the present era, respectively. 

It is not unlikely that in many places copper was largely used 
during the Stone Ages and before the Bronze epoch, since it was 
only after the art of making fire had been discovered that it became 
possible to manufacture bronze, whilst native copper could be 
fashioned without the aid of heat. Metallic relics of the Bronze Age, 



in the form of arms, ornaments, and domestic implements have been 
found in widely distributed localities. 

The mention of copper occurs in the Hebrew Scriptures, the metal 
being termed Nehosheih, from the root Nahdsh, to glisten. This A\as 
translated as x"^'^^^^' (chalcos) in the Septuagint, and Aes in the 
Vulgate ; the Greeks and Romans using the terms, however, both 
for copper and for the alloys brass and bronze. 

According to Pliny, the Roman supply was derived chiefly from 
Cyprus, and the metal thus came to be known a^s Aes Cyprium, which 
was gradually shortened to Cyprium, a name afterwards corrupted 
to Cuprum, from which are derived our modern terms Copper, the 
German Kupfer, and the French Cuivre. 

Copper was well known to the alchemists, and inasmuch as it 
was largely obtained from Cypi-us, an island dedicated to Venus, it 
was considered to be the metal specially sacred to the Goddess, and 
was generally known by that name in their writings, and symbolised 
by the sign o. The production of metallic copper on iron by the 
action of certain liquors from the Hungarian mines and other localities, 
was likewise known to the alchemists, and was a constant source of 
inspiration to them ; the changes were regarded for some hundreds 
of years as examples of the transmutation of the elements, until 
J3oyle showed that it was necessary to introduce copper into such 
solutions before that metal could be precipitated from them. 

The Development of the Copper Industry. — The mining and 
smelting of copper ores on a primitive scale have been carried on 
from time immemorial ; these operations were certainly practised in 
Greek and Roman days, and the deposits of Britain are said to 
have been known to the Phoenicians so far back as 1000 B.C. Percy 
refers to the finding of lumps of copper weighing 42 lbs., carrying a 
Roman inscription ; this metal was found in close proximity to mines 
in North Wales, which yielded an easily reducible ore, and he con- 
cluded that this was smelted in situ by the Romans. 

There are undoubted records of copper mining in this country 
in the time of Edward III., and in that of Elizabeth ; whilst the first 
authentic accounts of copper smelting date also from the latter period, 
relating to South Wales. It appears that one of the earhest establish- 
ments was situated at Neath — a fact recorded in a pubhcation of 1602. 
The works probably existed for a century before that date, and 
the copper smelters at Swansea were established about 120 years 

The processes employed for the primitive smelting of copper ores 
were, to a large extent, of the same nature as the crude operations 


practised generally for the extraction of metals in remote ages 
and by primitive races, as recorded from time to time by travellers 
and explorers. The furnace-hearth was a hole in the ground, working 
usually' on oxide ores with charcoal or Avood as fuel. This primitive 
furnace was later developed, by the addition of walls for enclosing 
the charge, until the '' shaft furnace " provided with an air blast of 
some kind was attained. The sulphide ores presented rather more 
difficulty in their treatment, but the production of metallic copper 
from sulphide materials by super-oxidation, in a process akin to the 
bessemerising of to-day. was developed in Japan centuries ago, and 
has been described by Professor Gowland. 

It would appear that during the middle ages, the art of reducing 
copper ores to metal on a comparatively large scale was practised 
simultaneously in Britain and in Central Europe ; first by primitive 
methods similar to those indicated above, developing later by succes- 
sive improvements into the employment of small blast furnaces. By 
about 1700, however, the methods diverged, and it is interesting to 
note that the different styles of working then introduced have 
persisted, until recent years, as the methods typical of these two parts 
of the world. In Wales, where the well-known furnace coal was one 
of the characteristics of the locality, as it still remains to-day, the 
smelting processes developed along the lines of reverberatory practice, 
for which such fuel is eminently suited, and this resulted in the 
estabhshing of the representative Welsh process. On the other hand, 
the enormous forests of Central Europe furnished wood suitable for the 
making of charcoal, a type of fuel which necessitates close proximity 
with the furnace charge, so that in these localities smelting was 
carried out in the shaft furnace, which gradually developed into the 
small blast furnace. At the present time, the solid fuel suitable for 
reverberatory practice is only obtainable in very small quantities 
in Central Europe, and the characteristic method employed there for 
copper smelting is that in which small blast furnaces are used, except 
that charcoal has been largely replaced by coke as the fuel. 

It is probable that the early ore furnaces of the primitive 
blast-furnace type in Britain were worked by Germans experienced 
in that class of work, just as at a later period in the history of the 
industry. Swansea coppermen were to be found in all parts of the 
world teaching other nations their art. Gowland reproduces a letter, 
dated January, 1583, protesting against the introduction of this 
foreign labour, whilst a second letter, dated July, 1585, which is also 
quoted, is of particular interest, as it gives evidence of a remarkable 
knowledge of the art of smelting, and, whilst illustrating an important 


feature of modern practice, indicates also the manner in which an 
astute smelterman was able to work profitably with difficult material 
so long ago as three and a quarter centuries. 
The letter is to the following effect : — 

" Ulricke Frosse to Robert Denham. 4th July, 1585. 

" To his loving friend, Robert Denham. 

*' Friend Denham, — I have me heartily commended unto you, you shall understand it 
we did lack ore more than 14 days ago, for we have found out a way to smelt 24 cwts. of 
ore every day with one furnace, the Lord be thanked, and if we may have ore enough from 
your side we may, with God's help, melt with two furnaces in 40 weeks 560 tons of ore, having 
reasonable jDrovision made for it, desiring you from hence-forward to send such ores as you 
have with as much speed as maybe, not caring what ore it is. Your ore of St. Dines is very 
hard to melt it, hoping we will overcome it what St. Ust ores will do, we long to see it. 

" This I rest, the Lord send you good success with your mines. And so I commit you 
to God. From Neath, the 4th of July, 1585. 

" Your friend, 

" Ulricke Frosse. 

" When you do send any more ore, if you can, send of all sorts, the better it will melt 
and with more profit." 

The sound principle of obtaining, when possible, one class of copper 
ore for the purpose of fluxing off the gangue from ore of another 
class, was thus recognised as a profitable feature of practice from 
comparatively early times. 

Copper mining and smelting in Staffordshire dates back a con- 
siderable time, certainly prior to 1686 ; the mines were situated at 
Ecton, and the smelter was at Elleston, near Ashbourne, where small 
blast furnaces were employed. Copper smelting in Lancashire, which 
is nowadays conducted on a comparatively extensive scale, appears 
to have commenced in 1720 with Cornish ores and smaller importa- 
tions from the West Indian and American Colonies. During the 18th 
century, the chief supply of the world's copper ore came from the 
Cornish mines, which even at that time, were deep and extensive. 
It seems, however, that for some peculiar reason, the Cornishmen 
were unable to smelt these ores with profit, nor indeed, to do more 
with them than to send the material to South Wales to be treated. 
There are numerous explanations for their failure, which have been 
discussed exhaustively by Percy. 

The centre of the copper smelting industry thus came to be located 
in the South Wales (Swansea) district, where circumstances were very 
favourable. The study of local conditions is one of great importance 
for metallurgists, and since this case affords a good example, it will 


be of value to refer briefly to those circumstances which rendered the 
Swansea district such an excellent centre for the industry. 

The extensive collieries in the locality rendered available an abund- 
ant supply of suitable fuel at a low price, and many of the smelters 
held a financial interest in them. The large coal was profitably used 
for home consumption or export, and the small, which, though dirty, 
still gave the long flame required, was very suitable for smelting work, 
and was reserved for that purpose. 

Further, Swansea w^as an excellent seaport, situated at a short 
distance only from Cornwall, the chief source of ore, and was also 
readily accessible to vessels carrying cupriferous ores and products 
from South America, Australia, and other parts of the world. This 
was a great advantage, in that the Swansea copper smelters, having 
a large variety of ores at their disposal, some with basic gangue, others 
^\ ith siliceous gangue, were in a position to make up furnace charges 
which were more or less self-fluxing, and thus avoided the necessity 
for purchasing and using barren fluxes. The finished products were 
also in a most convenient centre for distribution, at the seaport 
of Swansea. 

At the end of the 18th century, Great Britain was producing 
75 per cent, of the world's copper, the Cornish mines supplying most 
of the copper ore, and the Swansea smelters extracting most of the 
world's supply of metal. Stevens has summarised the position for 
1799, showing that '' from the Cornish ores 4,923 tons of refined copper 
were produced, and from the Welsh ores of Anglesea 2,000 tons. The 
great Mansfeld mine in Germany produced only 372 tons in that 
year, Spain's output was insignificant, and in the United States 
only a few tons were made. Russia and eJapan probably ranked next 
to Great Britain as producers, small amounts of ore from Austria, 
Scandinavia, and Italy made up the remainder. Thus at the com- 
mencement of the 19th century, the copper resources of the United 
States, Spain, Chili, Mexico, Australia, Tasmania, Canada, and South 
Africa, which now supply over 90 per cent, of the world's metal, were 
either undeveloped, or only yielded a few tons each ; Great Britain, 
which produced nearly 7,000 tons of copper at that time, extracted 
from its own ore supplies, a hundred years later, only 550 tons." 

It will be remembered that it was in connection with the develop- 
ment of Cornish copper mining that the use of steam power in engineer- 
ing was introduced and successfully worked out. On account of the 
increasing depth and extension of the Cornwall mines, the problem of 
disposing of the underground water became urgent, and led to the 
introduction of steam engines for driving the pumps, the Newcomen 


engine being installed on the Wheal Fortune Mine in 1720. The 
success of this engine led to increase both in depth and in extent of 
the workings, until it became impossible to cope with the pumping 
requirements by this means. At the right moment Watt brought out 
the modern steam engine, and the first Watt engine was erected in 
1777 at Chasewater, in Cornwall. It was the introduction of these 
improved methods of pumping which have made possible the success- 
ful development of present-day mining. Not only has the steam 
engine thus led to an increase in the supply of copper, by enabling 
the opening up of vaster deposits to be undertaken, but the develop- 
ment of engineering science which it has brought about, has caused 
a further consumption of the increasing quantity of copper which it 
has helped to render available for use. 

During the first half of the 19th century Great Britain retained 
its position as the chief copper producer of the world, and the 
Swansea smelters possessed advantages such as have been rarely 
enjoyed by any other body of manufacturers. They were able to 
impose what conditions they pleased on the producers and sellers 
of copper ore, as well as on the consumers of the metal, and as business 
men, were not slow to avail themselves of their opportunities to the 
greatest possible extent, strengthening their position by the formation 
of a combination known as the Associated Copper Smelters of Swansea, 
which controlled the price of the metal from 1850 to 1860. Percy 
gives an interesting account of the terms imposed by them under 
the name of returning charges, etc., as well as of the conditions of 
sampling, analysis, and sale, which were strongly in their favour. 

During these years of monopoly, the smelters were, on the whole, 
conservative in tendency from the metallurgical point of view, and 
few great developments in either processes or methods were devised : 
nevertheless, they enjoyed great prosperity, and their business attained 
such dimensions that Swansea remains one of the greatest centres of 
smelting industry in the world. The Welsh smeltermen had, more- 
over, acquired such proficiency in furnace management, and such 
knowledge of the working and control of copper charges, that their 
reputation had spread to all quarters of the world. 

Though from 1840 onward, the British copper mining industry 
commenced to decline, still for 20 years longer the Swansea smelting 
works prospered more and more as new mines were being opened 
abroad and thus furnished a constantly increasing supply of rich 
copper ore, cheap to purchase and easy to smelt. 

It was this development of foreign copper resources, and the 
unsatisfactory conditions which the producers received at the hands 


of the smelters, which was the cause of the eventual displacement of 
Swansea from its position as the leading seat of copper manufacture. 

In 1830. the production of copper ore in Chili had commenced 
and developed rapidly, Chili soon becoming one of the chief suppliers 
of ore to the Welsh smelters, whose independent attitude led to the 
first introduction of the copper-smelting industry on any large scale 
in America. Owing to the sailing conditions of the time, the simul- 
taneous coming into port of several ships laden with ore, instead of 
their arrival at regular intervals, enabled purchases to be made by the 
smelters at a remarkably low figure, tlie standard price of the metal 
being subsequently raised. Mine-owners commenced to seek for a 
remedy, their ultimate endeavour being to substitute, for the exporta- 
tion of their ores, smelting operations at or near the mines themselves. 
In 1 842 Lambert introduced reverberatory furnaces into Chili, and so 
great was his success, that in a short time they were in use throughout 
that country. In 1857 he erected the first blast furnace in Chili, 
and the smelting industry thereupon grew so rapidly that, whilst 
from 1856 to 1865 the copper exports from Chili were in the propor- 
tions of ore 21 per cent., regulus 38 per cent., and bars 40 per cent., 
they subsequently became ore IJ per cent., regulus 3| per cent., and 
bars 95 per cent. The ultimate effect was a widening of the market 
for the finished Chilian product, so that Continental purchasers were 
enabled to obtain their supplies of metal direct, instead of being 
obhged to purchase from the Welsh smelters on the unsatisfactory 
terms then prevalent. 

In 1842 the first large copper mines of Australia (Kapunda and 
later Burra Burra) were discovered, but developed slowly ; and in 
1844 the first copper mines of the Lake Superior district began work 
— on oxide ore. not on native metal. 

In 1850 an enormous development in the Chilian mines commenced, 
half the world's copper being produced from this source ; in 1859-60 
the Spanish mines at St. Domingo (Mason and Barry) were re-opened, 
as well as the Portuguese mine, the Tharsis. These mines were in 
reality operated in order to supply the wants of the sulphuric acid 
industry, the ore residues being subsequently smelted for copper at 
Swaasea. In 1862, however, the Henderson wet process for copper 
was introduced, for which these materials were very suitable, and the 
Spanish and Portuguese supplies became of considerable importance. 
soon afterwards coming under the control of a Scottish company. 

The competition from these new and abundant supplies of rich 
ores from Chili, Spain, and Portugal severely injured the production 
from the British mines ; increasing supplies led to a fall in the price, 



and one native mine after another shut down, the British supj)ly 
diminishing with considerable rapidity. 

In 1866 the great Calumet and Hecla mine at Lake Superior com- 
menced operations, and speedily became one of the most important 
sources of copper in the world ; the Moonta and Wallaroo mines in 
Australia opened about the same time, and in 1873 the Arizona mines 
started producing. In 1876 the enormous Spanish mines at Rio 
Tinto were re-opened, and soon rendered available large quantities 
of ore. Later, the Tasmanian supplies entered the markets. 

In 1880 a remarkable development in copper mining occurred 
with the discovery of the Butte camp in Montana ; this is now the 
greatest producer in the world. 

The later extensions of the copper minmg industry occurred in 
Utah, Tennessee, and Queensland, whilst within recent years the 
most important work on a large scale has been commenced in 
Tanganyika, in Nevada, and in Siberia. 

The developments in the smelting industry in most of these 
localities have proceeded, until the last few years, on very similar lines. 
During the first periods following the opening up of mines and works, 
ore was shipped to the custom smelters, most often to Swansea ; 
where, in the early days, mam^ of those connected with the smelting 
works had some sort of financial interest in the foreign mines. Later, 
the ore underwent its first smelting to matte in the mining district 
itself, the matte product being then shipped East for treatment, 
thus saving much of the freight-charge on useless gangue, as well as 
smelters' heavy returning charges, etc. At a later period the 
smelting operation was carried to a still further stage in the mining 
district, crude blister copper only being sent to Swansea or elsewhere 
to be refined. 

Gradually, electrolytic refineries were established somewhat 
nearer to the mining districts, and in the natural course of 
events, and where local conditions are not prohibitive, the pro- 
bability is that the whole cycle of operations from mining to the 
production of refined market metal will be carried out at the great 
camps themselves. 

At present, however, this is not generally the case, since the 
conditions under which the enormous refineries in the Eastern States 
of New York, New Jersey, and in Baltimore, etc., operate, allow of 
the cheaper production of electrolytic copper at points nearer to 
the distributing markets. At Anaconda, indeed, the fully-equipped 
electrolytic plant was shut down, owing to the commercial conditions 
such as have just been indicated, having rendered the refining of 



the anode copper at the Eastern refineries more profitable than 
electrolytic treatment on the spot. 

The Chief Features in the Development of Modern Copper 
Smelting Practice. — In the early days of copper smelting, the 
reduction of the oxidised ores, which were then chiefly available, 
was not a problem of very great difficulty, although losses in slag 
were likely to be very high, and the operation generally wasteful. 
\Mien, however, mines became deeper and sulphide ores had to be 
smelted, the problem became rather more complicated. In the first 
stages of development, such ores were probably roasted until as much 
sulphur as possible had been driven off, leaving practically an oxide 
charge to be treated by the older reduction methods involving the 
attendant extravagance in fuel consumption and large losses of 
copper in the slag. 

From these crude and wasteful methods the Welsh process was 
gradually worked out, and it will ever rank as one of the finest examples 
of highly developed smelting practice in the history of metallurgy, 
particularly when the times and working conditions are borne in mind. 
The process having received such full- treatment from most of the 
common text-books, it is not proposed to review it in detail here, 
since, moreover, it has been largely superseded by more modern 

As will be explained later, copper smelting of sulphide ores is essen- 
tially a fractional oxidation — chiefly of iron and sulphur — followed 
by the slagging or elimination of extraneous constituents of the ore. 
The Welsh process embodied a series of roastings and slaggings which, 
though most admirably adjusted for a substantial concentration of 
the copper in each succeeding product, allowed of the formation of 
slags in the first stages which carried but comparatively little copper, 
on account of the low tenor of the matte ; whilst the slags in the later 
stages of the process, containing more copper on account of their 
association with higher grade matte, were made in such relatively 
small quantity that their re-treatment for the recovery of these values 
did not involve very much loss of efficiency in the furnace operations. 

Later modifications of the process were chiefly devised with the 
view to reducing the number of operations, by eliminating the suc- 
cessive roasting stages, for which purpose oxidised materials, such as 
roasted or oxidised ores, were added to the charge. 

The Best-Selecting process, and the Nichol and James process are 
likewise valuable and ingenious modifications of the Swansea method 
for special work. 

In general, however, up to 1880, there had taken place but little 


change in principle from the older methods of smelting. The chief 
improvements involved a slow change in furnace size, and progress in 
several details in practice. The more important of these advances were — 

(a) In Roasting Practice. — 1865. Introduction of the mechanically 
driven furnace (the Briickner cylinder) ; not, however, adopted for 
copper smelting till many years afterwards. Later — Arrangements 
for using roaster gases for sulphuric acid manufacture. 

(b) In Reverheratory Furnace Smelting. — 1861. Gas firing intro- 
duced, but with very little success for copper smelting, even at 
the present day, 

(c) In BJast-Furnace Smelting. — Several very important changes 
were introduced in the construction of furnaces. 

1863. Elongation of the furnace. 

Rachette in Germany introduced the elliptical blast furnace. 
(Intended first for lead smelting ; rapidly adopted for copper matte 

1875. The water- jacketing of blast furnaces. 

The Piltz water-jacketed furnace was likewise first employed in 
lead smelting, and subsequently introduced into copper smelting 
practice. The principle had, indeed, been utilised in certain branches 
of iron smelting before this date, but for non-ferrous work the idea 
was new. 

Although the method of water-jacketing was recognised as 
leading to great improvement in the working of the furnace, its use 
was at first somewhat restricted, owing to various practical diffi- 
culties, and the ultimate great success was effected when in American 
practice, the plan of working the two principles of elongated furnaces 
and water- jacketing in conjunction, was adopted 

Commencing from 1880, and onwards, however, when production 
in the Far West began, enormous advances have been made, both in 
connection with the principles of working as well as in practical 
operation. These include — 

(1) Enormous increase in the size and capacity of furnaces of 

both the reverheratory and the blast-furnace type. 

(2) The application of the Bessemer process to copper mattes. 

(3) The development of the pyritic smelting principle. 

(4) The adoption of electrolytic refining. 

(5) The use of mechanically rabbled roaster furnaces. 

(B) The manufacture of sulphuric acid from blast-furnace gases. 
(7) The blast-roasting and sintering of sulphide fines. 


With an increased output of ore from tlie mines, and with increased 
consumption, stimulated by the growth of the electrical industry, 
the demand for metal increased so quickly that developments natur- 
ally^ followed with a view to an augmented and rapid production by 
more efficient and scientific processes ; especially since increased 
competition and poorer ore supplies necessitated a very decided 
lowering of the costs of production. To meet the enormous present- 
day demand for metal with the older methods and furnaces would 
have been impossible. The greatest stimulus to the adoption of 
these new or modified processes was the shifting of the chief producing 
centres from the older and more conservative influences to districts 
like the then newly awakening West, where, Avith ever-increasing — 
almost limitless — supplies of ore available, and free from the necessity 
of considering the capital invested in old plants, the men in charge 
of the work, untrammelled by old smelting customs which might 
stand in the way of rapid progress, were in a position to develop their 
ideas with originality and vigour. 

There may, nevertheless, be recalled the important share which 
British, and especially Swansea, workmen had in this great develop- 
ment of the industry. At many of the greater smelters in these new 
districts, Welsh furnacemen are still to be found, and large numbers 
went abroad in former days to take charge of such work, especially 
during the critical early stages. The principles underlying these 
modem improvements were, in many cases, first worked out by 
scientists in Europe. 

The Price and Cost of Production of Copper^ — The price 
of copper has been influenced to an enormous extent by financial 
speculation, so that until recent times it has fluctuated very con- 
siderably from year to year, the curve in fig. 1 relating to Best-Select 
copper, indicating this variation over a considerable period. The price 
of the other qualities of commercial copper follows this line fairly 
closely, electrolytic copper being from £2 to £4 per ton lower, and 
standard copper £3 to £6 per ton. The average value of the standard 
refined metal at the present time (December, 1911) is about £56 per ton 
in London, and about 12 cents per pound in New York. 

On three occasions during the past century, and once at least 
during the past decade, the market price of copper has been directly 
affected by more or less artificial conditions consequent on financial 
manipulation. The first of these instances was the 1860-1860 period, 
when the Welsh smelters held the monopoly of the copper trade, and 
were in a position to fix their own price ; the second was during the 
French combination of Secretan during 1887-9, which, as a result 










Fig. 1. — Fluctuations in the Price of Best Select Copper. 

of mere market speculation, caused fluctuations of price which 
amounted on one occasion to no less than £35 per ton within 



twenty four hours. The third instance was created by the American 

In 1899 the Amalgamated Copper Company was formed in the 
United States. This corporation was estabhshed in view of the 
enormously increasing production of the West, and of the extensive 
development of electrical industry which involved a greatly increased 
consumption of copper ; and it was probably designed to control the 
world's copper industry. Prices were raised gradually for some time, 
but in 1901 the Trust, as then constituted, failed, owing largely to 
trade depression in Europe. Heavy losses resulted, as well as ex- 
pensive law suits, and the price of the metal dropped again with great 
rapidity. Trade subsequently revived and expanded, the consump- 
tion of copper increased and appeared to overtake the rate of produc- 
tion, whilst stocks diminished and the price advanced, mitil, in 1907, 
copper was sold at well over £100 per ton. The American financial 
panic in the autumn of that year again reduced prices to a compara- 
tively low figure, and they have, on the whole, remained fairly steady 
since, though showing a tendency to decrease. Production has, 
meanwhile, increased very largely, and a steady price of 12 to 13| 
cents per pound yields handsome profits to most of the larger 
concerns. The present situation in the copper market is such that 
the enhanced production has again resulted in an accumulation of 
stocks, which has occasioned restricted output on the part of many of 
the principal smelters until briskness of trade development shall call 
forth increased consumption and more satisfactory prices. 

The question of price is one involving certain considerations to 
which attention may be drawn. The present conditions and the 
comparative steadiness in the copper market have been shown in a 
recent review to result in part from : — 

(1) The concentration of the copper industry in a few strong 
hands, which, whilst maintaining healthy competition, keeps the 
market free from such outside pressure as would reduce the price 
too much, and by restricting unprofitable output, brings production 
and consumption into equilibrium, making for stability. 

(2) The comparative cheapness of money, which has allowed of 
the financing for large production, with the prospect of absorption 
not being long delayed. 

At the same time, some of the richer and more cheaply worked 
mines of former times are gradually approaching exhaustion — recent 
instances of this will be readily recalled, whilst the disadvantages 
of having to work lower-grade deposits at greater depth have also 
tended to increase the price of metal. These conditions, on the other 



hand, have been counterbalanced by improvements in the mining and 
metallurgical processes concerned, by the opening up of new districts, 
and by the economies resulting from amalgamation of interests, 
involving closer organisation and enormous outputs of material. 

Apart from finance, two of the factors most likely to affect the 
price of the metal considerably are the possible replacement of copper 
for electrical transmission purposes by conductors of other metals ; 
and further, the enormous prospective production in the newer 
districts, such as Utah, Nevada, and Tanganyika, in the course of 
a few years. 

The cost of production of the metal is so dependent on local and 
general circumstances as not to admit of analysis in this place. 
Questions of locality, transport facilities, proximity to supplies of 
every kind, problems of labour, capitalisation, bye-products, and 
numerous similar considerations have such an important bearing on 
each individual case as to convey a definite meaning only to the 
man on the spot. In the same way, detail costs of each stage of the 
copper smelting processes are influenced by similar circumstances. 

Broadly speaking, the average total cost of production and market- 
ing at present may be taken as being somewhere about 10 cents per 
pound of copper ; in certain specially favoured cases, 9, 8, or even 
7 cents per pound. The newly opened low-grade "porphyry'' 
camps at Utah and elsewhere, which have been commenced under an 
enormous capitalisation, anticipate a production at a cost of about 
6 cents per pound when steady and normal running is in progress. 

A recent analysis gives interesting information Avith regard to the 
cost of production estimated at different plants. Of the American 
output of about 480,000 tons in 1909— 

Almost 3-5 per cent, was produced at a cost of 7-14 cents per lb. (Nevada). 

1-8 „ „ „ 7-98 „ (Baltic, Superior). 

10-5 „ „ „ 8-9 „ (Utah, etc.). 

48-3 „ „ „ 9-10 ,, (Boston and Montana, 

Calumet and Hecla, 
9-0 „ „ „ 10-11 „ (Utah Consolidated, 

Tennessee, etc.). 
20'0 „ „ „ 11-12 ,, (Anaconda, Arizona, 

1-8 „ „ „ 12-13 

M M „ „ 13-14 

1-4 „ „ „ 14-15 „ (Tamarack). 

M ,, „ „ 15-16 

M „ ,, ,, 16-17 

0-1 „ „ „ 1709 



Copper Statistics- — The outstanding features which attract 
attention in the statistics of copper production Avill be most readily 
seen from the curves of fig. 2. The enormous increase within recent 
years in the total output of metal, and the overwhelming proportion 
produced by the United States of America, is clearly indicated. 
The curves also show the practical extinction of the native supply of 
Great Britain and the steady output of Spain and Germany. 

An analysis of the total production for the year 1910 is given in the 
following Table T. : — 

TABLE I.— The Production of Copper (Short Tons of 2,000 lbs.). 



United States of America, . . 549,114 


Canada, . 

. 26,998 


1909. 1910. 




►North America, . 644,058 645,927 

Mexico, . 

. 63,085 


Cuba, . 






Bolivia, . 



-South America, . 60,911 63,101 



20,502 > 

Spain and Portugal, 

. 58,447 





Russia, . 



Norway, . 



Sweden, . 




.Europe, . . 127,283 135,788 

Italy, . . . 



Austria, . 



Turkey, . 



Great Britain, . 






Africa, . 






Total, .... 



In Table II. is indicated the distribution of the American pro- 
duction among the various States. 







—North American Production of Copper 

(in Short Tons of 2,000 lbs.). 










California, . 












Michigan, . 




Montana, . 






31,944 (about 6,000 tons in 1908) 

New Mexico, 







62,521 (about 35,000 tons in 1908) 

Wymoing, . 



South and East, 




Other States, 








There will be noticed a decline in the production of the United 
States during the year 1910, resulting from the present movement to 
restrict output whilst the large accumulated stocks of metal are being 
absorbed. The movement is probably more or less temporary, and 
is being largely directed by American financiers who are endeavouring 
to bring about an international agreement on the subject. 

Regarding the American output, the marked movement for cur- 
tailment in Montana has reduced the output of that State to such 
an extent, that the position it gained in 1909, of being the greatest 
producing State once more reverts to Arizona. The increases from 
Nevada and Utah, in which developments on a large scale are 
commencing, may be noted. 

Percy, John, 
Growland, William, 

Stevens, H. J., . 
Brown, N., and Turnbull, C. C, 
Enz/inefring and Mining Jom-nal, 
Mineral Statistics of the United Kingdom 
Mineral IndiLstry. 


" Metallurgy (Copper)." 

Presidential Address, Trans. Inst. Mining and 

Metallurgy, vol. xvi., 1906-7, pp. 265-291. 
" The Copper Handbook." 
" A Century of Copper." 
" Copper Production." May 6th, 1910, p. 891. 



The Uses of Copper: as lYIetal and as Alloy— The Physical 
Properties of Copper— Effects of Impurities— 
IVIechanical Properties— Chemical Properties. 

The Uses of Copper, — Generally speaking, the industrial appli- 
cations of copper involve its employment in two forms : — 

(1) As metal. (2) As a constituent of alloys. 

The more limited use in the form of copper salts is of chemical 
rather than of metallurgical interest. 

Copper in the metallic form is employed for three classes of 
work : — 

(a) For electrical purposes. 
(h) For engineering purposes. 
(c) General industrial uses. 

(a) Electrical Uses. — Of late years the marked growth in the 
consumption of copper has arisen very largely from its usefulness as 
a conductor of electricity ; the increased demand for the metal with 
the development of electrical enterprise being a well-marked feature 
in industrial progress. It is estimated that from 60 to 70 per cent, 
of all the copper produced is utilised for this purpose, and metal is 
specially prepared and sold under the designation of " high-conduc- 
tivity copper/' The demand has, to a large extent, increased irre- 
spective of price up to recent years, owing to the necessity of employing 
copper for such purposes, though the natural economic factor that 
an enhanced price of the metal tends to some discouragement of 
expansion and of fresh electrical enterprise, has exerted considerable 
effect in checking consumption. 

It is merely necessary to enumerate some few^ of the present aspects 
of electrical industry in order to realise the enormous absorption of 
copper in this connection, as, for instance, electrical traction, lighting, 
and power, the telegraph, and the telephone. With reference to the 
use of the metal for this work, it is important that certain mechanical 
as well as electrical requirements should be fulfilled, for in many | 
branches, considerable strength of the material is also requisite. ^ 



The demand of the electrical engmeer is that as a conductor, the 
copper shall offer a minimum of resistance to the passage of the 
current, and for this requirement the metal must be in a condition 
of ver}' great puritA'. With but few exceptions, this necessitates the 
purification of the copper by electro-deposition. Electro-deposited 
metal as produced at the refineries is, however, not immediately 
suitable for drawing into Avire, owing to the weakness and porosity 
inherent in the material prepared by this method. It must, therefore, 
be melted, brought to pitch, cast into bars, and these bars trans- 
formed into wire, which operations require to be conducted with 
much care in order to keep the metal in as pure a condition as 
possible for its work. It may be noted that within recent years, 
several processes, notably those of Cowper-Coles and Elmore, have 
been put into operation for the direct manufacture for electrical 
purposes, of electrolytic-copper wire of the requisite strength. 

The mechanical qualities demanded of the metal for such purposes 
as telegraph work may be indicated by the two specifications of wire 
for the British Post Office, which are appended : — 

A. Post Office Specification. 

Weight, loO lbs. per mile. 
Minimum diameter, -95^". 
Maximum diameter, -98".* 
Minimum breaking strain, 490 lbs. 
Minimum number of twists, 25 in 3 inches 

B. Post Office Specification. 

Weight, 500 lbs. per mile. 
Minimum diameter, -ISoi". 
Maximum diameter, 1381'. 
Minimum breaking strain, 950 lbs. 
Minimum number of twists, 30 in 6 inches. 

Wraps required, <5 times round wire of its 1 Wraps required, G times round wire of its 

'\vn diameter, unwrapped, and again 
wrappcfl without breaking. 
Maximum resistance per mile at 60° F., 

5-857 (t>. 

own diameter, unwrapped, and again 
wrapped without breaking. 
Maximum resistance per mile at 00° F. 

2-928 a,. 

The following figures afford some indication of the increasing 
demand for copper in two branches only of electrical industry : — 

1902. 1907. 

Mileage of wire existing for telegraph purposes, . 3-9 million miles, 53 million miles. 

„ „ telephone purposes, 10-9 „ 28-2 „ 

(b) Eiigineering Uses. — Metallic copper finds application in marine 
shipbuilding and engine work, as well as in railway and locomotive 
work, where the metal is particularly employed for steam pipes, 
and for fire-box plates and stays, sometimes also for boiler tubes, on 
account of its high conductivity for heat, combined with toughness. 
The questions of suitable composition, and the other requirements 

* This figure indicates thousandths of an inch. 



of the metal intended for these purposes, has been a subject for 
discussion by some of the leading marine and locomotive engineers. 
Useful information on the subject will be found in the reports of some 
of these discussions at the Institution of Mechanical Engineers. 

The following tests are required for copper plate (best quality) 
intended for locomotive fire-boxes on the Lancashire and Yorkshire 
Railway, taken from standard specifications given by their Chief 
Mechanical Engineer at the Institute of Metals :- 

Bending Test. — Pieces of the plate shall be tested both cold and 
at a red heat by being doubled over on themselves — that is, bent 
through an angle of 180° — without showing either crack or flaw on the 
outside of the bend. 

Flanging. — Plates must not show any defects in flanging. 

Tensile Test. — Ultimate breaking load, 14 tons per square inch ; 
Elongation, 35 per cent, in 8 inches. 

Analytical Test. — To be made at contractor's expense. 

The copper upon analysis to give the following results : — Arsenic, 
not less than 0-35 per cent, nor more than 0-55 per cent. ; other 
foreign elements, exclusive of combined oxygen, not to exceed 0-25 
per cent 

Clauses are also inserted as to stamping, inspection, and the giving 
of testing facilities. 

Typical analysis of such plates show — 

Copper, .... 99-30 per cent. 
Arsenic, . . . . 0-43 to 0-5] 

Oxygen, .... 0-1 per cent. 

Impurities, chiefly antimony, lead, iron, nickel, tin, and sulphur, 
not exceeding 0-25 per cent. 

The average test on a number of plates gave — 

Tensile strength, . . 14-66 tons per square inch. 

Elongation on 8 inches, 
Contraction of area, . 
Close bend test, 

The effect of temperature and 

43-36 per cent, of original length.] 

45-9 per cent. 


the influence of impurities on the 

mechanical properties of the metal intended for engineering purposes 
are of very great importance, and much attention has been devoted 
to researches in this subject, particularly by Milton and Le Chatelier, 
whose published experience gives important information of much 
practical value. The main conclusions arrived at from practice have 


had reference to the general effects of impurities in hardening the 
metal, and the general tendency of heat to soften it and to increase 
the dnctility. The diverse effects of different impurities on strength 
and ductihty will be reviewed in detail at a later stage. 

(c) General Industrial Uses. — Copper as metal is also employed 
to a considerable extent in certain important industries, as in textile 
manufacture, where it is used for the rollers in calico-printing ; and 
it is in general industrial use in the form of copper heaters, vats, coils, 
pans, and the hke. and occasionally also for roofing and sheathing. 

Uses of Copper Alloys. — Between 20 and 30 per cent, of the 
copper produced is employed in the form of alloys. The more 
important of these are : — 

Brasses ; alloys of copper and zinc. 
Bronzes ; chiefly alloys of copper and tin. 
Coinag'e Alloys ; of gold and silver with copper. 
German Silver ; alloys of copper, nickel, and zinc. 
Special Bronzes ; alloys of copper with such metals as aluminium 
and manganese. 

It is further not unlikely that several classes of ternary alloys, at 
present still under investigation, may have important industrial 
application in the future. Among such alloys may be mentioned 
the copper-aluminium series alloyed with other metals, Monel metal 
and the Monel steel series, etc. 

Of the above alloys, the brasses are by far the most widely used. 
It may be recalled that the advantages possessed by alloys of copper 
and zinc are in large measure due to their increased strength and 
hardness ; to the fact that they are more fusible, and more fluid when 
melted, and so give good castings ; that they are characterised by a 
good colour and high lustre, as well as by the factor of cheapness 
resulting from the addition of a less costly metal — zinc — in their 

The uses of the copper alloys may also be arranged in two classes — 
{a) engineering uses, and (b) general uses. Of the brasses, those con- 
taining upwards of 70 per cent, of copper may be rolled cold, whilst 
the alloys with less than 70 per cent, are hot-rolled. 

In the engineering industry large quantities of 70/30 brass are 
utihsed in the form of condenser tubes, whilst for the multifarious 
requirements of general engineering work, very considerable amounts 
of brass of lower tenor are employed in the forms of taps, pipes, fittings, 

Muntz metal, the 60/40 brass, finds extended application for the 



sheathing of ships, whilst the employment of brass and of the other 
alloys for all manner of articles of general utility is a matter of common 

The close connection between properties, constitution, and the 
equilibrium diagram of these various classes of alloys has become 
manifest to a marked degree within recent years, and the effects of 
thermal treatment partly in modifying their constitution, and thereby 
the properties, and also in controlling the condition and distribution 
of the constituents, are at the present time having an important 
bearing on the manipulation of these alloys in the industries manu- 
facturing them and adapting them for their various uses. The study 
and application of these equilibrium diagrams are highly important to 
those who have to deal with these alloys on an industrial scale. 

60 50. 40 30 20 

I'ig. 3. — Equilibrium Diagram, Qu-Zn Series. 

The Properties of Copper. — The properties of the metal which 
render it of such service in the arts and industries are mainly its 
high electrical conductivity, its great ductility, malleability, and 
toughness, which enable it to be readily worked up into the different 
forms in which it is employed, its high thermal conductivity, and its 
resistance to the various agencies which lead to corrosion. These are 
consequently the properties to which close study is directed. Of per- 
haps still greater importance is a knoAvledge of the influence exerted 
upon these properties by the circumstances Avhich usually attend 
working practice ; such as, for example, the various common im- 
purities, and the variations of temperature, as well as the previous 
mechanical and thermal treatment. These can only be indicated in 
general terms here, references to authorities on the different branches 
being given later. 



Physical Properties. — The colour of copper is familiar, being a 
fine salmon pink. The appearance of the fractured surface is a useful 
guide in several respects as to the condition of the metal, and in the 



OF Impurities on the Electrical Conductivity. 


Addicks. j Johnson, 















! . 




Pure copper, 



\ Copper with — 

Copper with — 

1 Manganese, 





98-6 ! 0-01 



66-8 0-02 


1 Nickel, . 






Oxygetj, . 




























Platinum, . 

















Bismuth, . 












Cadmium, . 











Sulphur, . 





Cobalt, . 


































- - 

Tin, . 




































process of manufacture the refiner relies upon this appearance as an 
important criterion of the progress of the refining operation. Copper 
containing an excess of oxygen, for example, has a purpUsh-red colour 
and a coarse brick-like fracture ; this is known as " dry copper," and 
the metal is brittle and commercially useless when in that form. 
The ingot of dry copper is also characterised by a depression running 
along the surface. Tough copper (" tough-pitch '') the mechanically 
useful variety resulting from the furnace-refining operation, possesses 
a bright salmon-coloured fracture, finely granular to silky in appear- 
ance, whilst " overpoled copper " also brittle and industrially valueless 
whilst in that condition, has a very light salmon-coloured fracture, and 
is more coarsely fibrous. 

The melting point of copper is 1,083° C., and is slightly lowered by 
the small quantities of impurity usually present in commercial metal. 
Molten copper is of a pale apple-green colour. The boiling point 
under ordinary conditions is about 2,300° C. (1,700° C. in vacuo). 
The electrical conductivity is of much importance. Copper ranks 
second only to silver as a conductor, the relative conductivity of 
the best copper being about 98 compared with silver as 100. The 
resistance of 12 inches of pure copper wire, 0-001 inch in diameter, 
is 9-612 ohms. The conductivity of the metal is decreased by 
mechanical working, and it follows the general straight -line law 
connecting conductivity and temperature. 

The effect of even small quantities of impurity on this property 
is very marked, so much so that only the purest varieties are suitable 
for electrical work, and for this reason electrolytic refining is often 
a necessary operation in the manufacture of copper intended for 
this purpose. 

Table III. on preceding page, summarises the results of the work of 
Addicks and Johnson, and indicates the effects of small amounts of 
different impurities on the conductivity of the metal. 

The notoriously destructive effect of arsenic on the conductivity is 
very apparent. 

The influence of most of the common impurities is of a similar 
nature, and detailed investigations indicate that the effect is more 
or less progressive as the quantity increases — within the limits usually 
present in commercial metal. The results of Hiorns and Lamb's 
experiments with reference to arsenic and antimony are indicated in 
Fig. 4. 

The specific gravity of copper naturally varies according to its 
condition and composition. When pure and in the worked state, its 
density is 8-95 ; cast metal, more open and inclined to porosity, has 



a density of about 8-2 to 8-6, depending on the purity, rate of cooling, 
etc. Impurities lower the specific gravity. 

The conductivity for heat of the metal is high, being 898 compared 
with gold as 1,000, and as a conductor it is two and a-half times more 
efficient than iron. It is this property, combined with its toughness 
and resistance to corrosion, etc., which largely determines its employ- 
ment for heaters, steam-coils, and the like. 



M 10 



>2 r^ °^ 






■S 20 










OS 10 1-5 20 2-5 aOperCent 
Proportion of Impurity 
I^ig- -i- — Influence of Arsenic and Antimony on the Electrical Conductivity of Copper. 

Poiver of Dissolving Gases. — When molten,, especially under re- 
ducing conditions, the metal possesses the property, common 
to many others, of absorbing gases such as carbon monoxide, 
hydrogen, hydrocarbons, sulphur dioxide, etc., which are more- 
over, to a large extent insoluble in the solid material, and 
are. therefore, often liberated at or about the moment of 
solidification ; though some may remain dissolved. This action is 
one of the causes of the difficulty which is experienced in making 
sound castings of the metal, particularly since the gases mentioned 
are present in quantity during the poling and refining operations. 
The presence of certain materials in the copper, as in the case of steel. 


appecars to reduce the dissolving power of the liquid metal for these 
gases, or possibly to increase their solubility when the copper is solidi- 
fying, and in this way tends to minimise their injurious effects. It 
would seem that one of the functions of the cuprous oxide, which is 
purposely introduced into the metal when " bringing it up to pitch,'' 
is to exert this action. The ridge in the ingot of overpoled copper is, 
to some extent, accounted for as being due to the effects of the evolved 
gases, and this appearance indicates the absence of the requisite 
quantity of cuprous oxide necessary to counteract the effect. 

Copper is also supposed to be capable of holding certain quantities 
of gas in solution after it has become solid, and the resulting metal 
is more brittle and often commercially useless. Several of the charac- 
teristics of overpoled copper probably arise from this cause also. 

Impurities * in Copper, — In view of the marked influence of 
impurities on the properties of metallic copper, it may be advisable 
in this place briefly to review the results of recent scientific work as 
to the condition in which they exist in the metal, thus offering some 
clearer indication of the manner in which they affect the mechanical 
and other properties. The common impurities in ordinary commercial 
metal may be oxygen, arsenic, antimony, bismuth, lead, and to 
smaller extents, iron, sulphur, tellurium, and selenium. 

A factor of much importance is that the effect of two or more of 
the common constituents when present together, may be of even 
greater moment than that of each one separately, and in 
this connection Hampers classical work should be consulted. The 
investigation of the joint effects of impurities becomes so complex 
that systematic study progresses but slowly. Metallographic work 
is, however, revealing much evidence, and the researches in progress 
at present at several laboratories will, when published, afford greatly 
increased knowledge on the subject. Recent papers by F. Johnson 
give valuable detailed information (see References, p. 34). The 
importance of oxygen in this connection is particularly marked : its 
effects are profound, since in addition to its own specific influence as 
oxide, it also brings about chemical changes in some of the other 
constituents, thus leading to the formation of entirely new com- 
pounds possessing quite different properties. The beneficial influence 
of certain definite proportions of oxygen in addition to the other 
constituents of commercial copper is well known in practice, and 
has been systematically studied by Hampe, and later by several 

* The term " impurity " might in several instances be replaced by the word " consti- 
tuent," since many so-called impurities are purposely added for conferring desired properties 
on the metal. It is here taken as implying elements other than copper. 


other workers with more delicate means of investigation at their 

Oxygen in Copper. — Molten copper has the power of dissolving its 
oxide, CuoO. When the melted metal is exposed to oxygen, this 
oxide is produced and passes into solution in the hquid, yielding a 
series of binary alloys, of which the oxide acts as the second con- 
stituent. The equilibrium diagram of the series, as worked out by 
Heyn * (see Fig. 5). affords a good indication of these relationships. 











*^^ / 




^ctic So/I 





Cap^us Oxtdt Per Cent 




0896 A/2 

Oxygen Per Cent. 

Fig. o. — Relations of Copper and Oxygen. 

and throws Ught on several features connected with the presence of 
oxygen in copper. 

It will be observed that w hen molten oxygenated metal containing 
less than about 0-38 per cent, of oxygen solidifies, copper crystal- 
lises out first, whilst later, in between the copper crystals, there soli- 
difies a eutectic of copper and cuprous oxide. This eutectic contains 
about 3-45 per cent, of cuprous oxide, equivalent to 038 per cent, of 

* The temperatures given in Heyn's diagram require revision in the light 6f later 
knowledge, and have been omitted here. 


oxygen ; it melts at a temperature about 18° C. below that of the pure 
metal. The presence of this material, which is of a blue colour 
when viewed under the microscope, constituting slightly more fusible, 
tough, non-conducting areas between the copper crystals, accounts 
for many of the well-known effects of oxygen in metallic copper. 

When oxygen is present in quantities above the eutectic proportion, 
the first constituent to solidify from the molten over-oxygenated 
copper is brittle copper oxide, and the presence of such brittle material 
disseminated through the metal explains why " dry copper " cannot 
be worked. 

The effects of comparatively small quantities of oxygen are greatly 
increased on account of the fact that one part of oxygen, when present 
as cuprous oxide, yields a constituent in almost nine times as great a 
proportion by weight alone, since CU.3O : :: 142 : 16 or 9 : 1 ; whilst 
oxygen existing as oxide - eutectic is represented in the ratio of 
nearly 30 : 1 . The presence of excess of copper oxide in the metal is 
particularly dangerous when copper is to undergo annealing in a 
reducing atmosphere, since the reducing gases acting upon the oxides 
at the crystal boundaries destroy them, thus tending to produce that 
rottenness in the material which is so often encountered under such 

The great value and importance of oxygen in copper lies in its 
property of bringing the metal up to pitch as indicated above. 

The effect of carbon on oxygenated copper was the subject of much 
enquiry in early years. It was thought at one time that the influence 
of carbon per se in the copper was responsible for the beneficial effects 
resulting from the melting of brittle '* dry " copper with carbon, but 
the work of Percy, since confirmed, showed that its sole action is in the 
reduction of the injurious excess of oxide. 

In addition to the specific influences of oxygen as just recorded, 
and to its important physical effects with regard to the solubility of 
gases, etc., oxygen in copper performs other valuable functions, by 
forming with reduced impurities which are exceedingly dangerous, 
oxygenated compounds more infusible and more insoluble ; and this 
has the effect of segregating or distributing such injurious impurities 
into forms and positions much less harmful. 

Arsenic in Copper. — When arsenic and copper are melted together 
chemical combination occurs, and a series of arsenides is produced ; the 
system, which has been investigated by Friedrich (from whose work 
the following diagram has been constructed), Hiorns, Bengough & Hill, 
and others, being one of considerable complexity. With proportions 
of arsenic such as are usually present in commercial coppers, the 


•' .■^ 

-■■ -■'■'•^•'»c.^o.-\--V 

■J ^ 

O !>• 

§ II 


Lh '^ 

a :: 







*' tough-pitch " copper. This tough copper generally contains certain 
impurities which render the metal exceedingly useful for mechanical 
service, and their presence is, indeed, almost essential in copper 
intended for such purposes. At the same time, such elements would 
render it absolutely unfit for the other uses just specified, where 
purity is practicalh- the first necessity. 

The standard works and the papers indicated in the appended list 
of references should be consulted for details concerning the effect of 
each circumstance on the several mechanical properties ; certain 
general considerations must, however, be noted here. 

Not only should the composition of the metal be carefully con- 
fsidered, but attention must be directed to the actual condition and 
distribution of each constituent. Owing largely to the difficulties of 
detennining the oxygen contents in copper, and to a want of definite 
knowledge as to the condition, amount, and effects of the dissolved 
gases in the metal, the information at present available is not suffi- 
ciently concise to alloAv of a systematised statement being made as to 
the direct influence of the constituents on the mechanical properties. 
This is more especially the case since the other attendant circum- 
stances of working practice may react through these to a considerable 

Many of the more general results have, however, long been known 
to engineers from practical workings and these have been placed on 
record from time to time. 

The malleability and ductility of copper are considerable. Cold 
rolling and hammering causes a reduction in this respect, and the 
metal is hardened, but the properties are restored by anneahng. 
The annealing effect commences at about 300° C, but proceeds more 
effectively at higher temperatures, the factors of annealing tempera- 
ture and duration necessary for annealing being inversely connected. 
The impurities which influence these properties most adversely are 
bismuth and tellurium. The effect of other constituents, oxygen per 
86, sulphur, and iron, in the quantities usually present in commercial 
copper, is very small. Arsenic and antimony up to 0-4 or 0-5 per cent. 
have no deleterious effect on the malleability and ductility of (•()[)[)er 
0/ the correct pitch, and may even improve the metal when tested in 
the cold ; the hot malleability is, however, somewhat decreased. 

The presence of impurities raises the temperature required to 
bring about the full effects of annealing after the metal has been 
hardened by mechanical work. This action is probably explained by 
the interference of the impurities upon the molecular freedom of the 
metal, which controls the mechanism of annealing. The conditions, 


whether reducing or oxidising, during annealing, may exert an im- 
portant influence on the results. 

Hardness. — Pure copper is a comparatively soft metal. It is 
hardened by mechanical work — the hardness of rolled copper, deter- 
mined by the Brinell Test, being 74 compared with mild steel as 100 
— and by the presence of even small quantities of impurities, tin 
possessing a particularly marked effect in this connection. The 
worked metal is softened on annealing. 

Tensile Strength and Elongation. — The strength of copper, being 
a property of such practical importance, has been the subject of much 
extended investigation. The work has, however, been conducted 
under such a great variety of conditions, many of which have been^ 
left unrecorded, that co-ordination of the results is barely possible, 
and does not allow of establishing on a definite basis the effect of 
different influences on this property of the metal. Later work, some 
already published, some still in progress, should eventually allow ofi 
more general standardisation than is at present possible. The tensile 
strength of pure cast copper is 8 to 9 tons per square inch. Mechani- 
cal work causes an increase in the value up to 14, or even 16 tons, 
cold work exerting a still more marked influence ; whilst 33 tons and 
more per square inch has been recorded with cold-drawn fine wire. 
The elongation varies according to the mechanical work which the 
metal has undergone ; the amount ranges from 35 to 40 per cent, and 
upwards, measured on a 3-inch length. 

Tensile strength is reduced on annealing, but never to so low a 
degree as that of the cast material, the usual figure being 12 to 14 
tons per square inch. The effect of temperature in reducing tensile 
strength, especially when impurities are present, is important from 
the industrial point of view. The reduction in strength caused by 
annealing appears to be considerably smaller in the presence of 
arsenic and antimony. 

Arsenic increases the tensile strength of copper when the metal 
is of the correct pitch, "generally to well over 15 or 16 tons, in the 
presence of the proportions usually found. Antimony has a similar 
effect. Some workers state that, within certain limits, the strengthen- 
ing effect of this element is even more pronounced. Excess of 
antimony exerts, however, a much more adverse influence than does 
excess of arsenic. The elongation is increased by the presence of 
moderate quantities of arsenic. 

Oxygen per se, when present in moderate quantity in copper, has 
but little effect on the tenacity. Bismuth, tellurium, sulphur, and 
lead are the impurities which lower the strength, even when present 


ill minute quantities, and especially on heating. Bismuth in the 
proportion of 0005 per cent, lowers the malleabihty and ductiHty 
considerably, and recent reports state that 002 per cent, bismuth 

I renders copper cold short, that 005 per cent, makes it red short, and 
that 0-005 per cent, is the limit for electrolytic copper which is to be 
rolled. The deleterious effects of bismuth are, as already explained, 
to some extent masked by the presence of arsenic and by oxygen. 

The strength is increased by the presence of nickel, tin, and zinc in 
the proportions usually present in the commercial metal ; these are, 
however, generally small. 

From the foregoing review, indications will be afforded of the 
reasons for the choice by engineers of '* tough-pitch " copper for 
much of their work, and the explanation for the 0-3 to 0-5 per cent. 
arsenic often particularly specified for. The frequent use of arsenical 
coppers for such purposes as fire-box plates will also be understood, 
since the arsenic not only improves the mechanical properties of the 
metal, but ensures the retention of rigidity and strength at the high 
working temperatures required, to a greater degree than would have 
been the case had pure copper been employed. 

The effect of the above factors on the elastic limit of copper, 
is also very marked and of much importance, the influence 
being closely analogous to that produced on the other mechanical 

Chemical Properties. — The atomic weight of copper is 63-57. 
The metal is unchanged in dry air at ordinary temperatures ; in the 
presence of moisture and of carbon dioxide a green coating of basic 
carbonate is produced. When heated in air, a black scale, consisting 
of cuprous oxide, Cu^O, is obtained, which is readily detached by 
quenching and hammering. Water at ordinary temperatures is with- 
out effect upon copper ; concentrated sulphuric and nitric acids have 
Uttle action upon it in the cold, but attack it on heating. The best 

* solvent for the metal is dilute nitric acid, which dissolves it very 
readily. Copper is liable to corrosion when subjected, whilst hot, to 
the action of chlorine or hydrochloric acid gas ; this action has pro- 
vided an explanation of the corrosion of copper boiler tubes where the 
coal employed had been exposed to sea water. 

^ Copper is deposited from solution as a dull red, spongy mass, 
by iron, zinc, or aluminium, but it is more electro-positive than gold 
or silver, and readily precipitates these metals from solutions of 
their salts, these effects being extensively made use of in practice. 
The metal possesses a powerful affinity for sulphur, and this property 
has very important applications in the smelting processes. 




Copper readily alloys with gold, silver, tin, zinc, and nickel, but 
not with lead or iron. 


Composition and Properties of Metal for Railway and Locomotive Work (p. 20). 

Proc. Inst. Mech. Eng., 1893 ; Dean, p. 139 ; Blount, p. 1(34 ; Watson, p. 168 ; Gowland 

p. 176; Aspinall, p. 193; Tomlinson, p. 182. 
Webb, F. W., "Locomotive Fire-box Stays." Proc. Inst. C.E., 1902. 
Milton, J. T., " The Treatment of Copper for Steam Pipes." Inst. Marine Eny., 1 908-9.^ 
Hughes, C, "Non-ferrous Metals in Railway Work. " J. Inst. Metals, Sept. 1911. 
Law, E. F , " Alloys." 
Influence of Impurities on the Electrical Conductivity of Copper (p. 24). 

Lawrence Addicks, Trans. Amer. Inst Elect. Eng., 1903, vol. xxii., pp. 695-702 
Electro-Chemical Industry, 1902-3, pp. 580-583; Trans. Amer. Inst. Min. Eng 
1906, vol. xxxvi., p. 18. 
Walker, A. L., Mineral Industry, 1898, vol. vii., p. 248. 
T. Johnson, "Some Features in the Metallurgy of Copper." Proc. B'ham. Met. iSoc. 


Hlorns and Lamb, "Influence of Arsenic and Antimony on Copper." Journ. Soc 
Chem. Ind., May, 1909. 
Condition and Influence of Impurities on the Mechanical Properties of Copper (p. 32). 

j Zeitschrift jiir Berg. Hutten and Sal. Wesen, 1873, xxi., 218 ; 1876, xxiv.. 2t^ 
ampej qj^^^^j^^^ Zeituiuj, 1892, No. 42, p. 16. 

f Reports, Royal Tech. Testing Institute. CharlotteiJ 
Heyn, E., " Copper and Oxygen." j burg, 1900, p. 315. 

\ Metallographist, vol. vi., 1902, p. 48. 
Arnold, Engineering, vol. Ixi., p. 176. Feb. 7, 1896. 
Roberts-Austen, Second Report, Alloys, Research Committee. Proc. Inst. Mech 

April, 1893, p. 114. 
Rudeloff, Mittheil. Konig. Tech. Versuchs. Anstalt., 1894, ii. (6), pp. 292-330; 

16«, pp. 171-219. 

Lawrie, Bull. Amer. Inst. Min. Eng., 1909, jip. 857-66. " Bismuth in Wire Bar Copper.'^ 
Johnson, F., " Impurities in Tough Pitch Copper containing Arsenic." Proc. Inst, of 

Metals, 1910, vol. iv.; No. 2, p. 163, et seq. 
Johnson, F., "The Influence of Impurities on the Properties of Copper." Mefrd- 
lurgical and Chemical Engineeritig, Oct. 1910, p. 570. 
"Annealing of Copper and Diseases of Copper." Ibid., Feb. 1911, p. 87. 
"Notes on the Metallurgy of Wrought Copper." Ibid., August, 1911, p. 396. 
See also Standard Specifications for Copper Wire-Bars (recommendations by the Committee 
of the American Society for Testing Materials). E7ig. and Min. Journ., Jan. 20, 1912, 
p. 181. 



Compounds of Copper— Copper IVIattes— The Varieties of 
Commercial Copper— Ores of Copper— Preliminary 
Treatment of Ores, Sampling-. 

Compounds of Copper, — From the smelting point of view, the 
three most important classes of copper compoiuids are the oxides, 
the sulphides, and the silicates. 

Copper Oxides. — Of the oxides, two are of importance — cuprous 
oxide, Cu.p, and cupric oxide, CuO — the first-named particularly, 
having extensive connection with smelting practice. 

Cuprous oxide is black when in the massive form, and has a red 
hematite colour when powdered. It is readily formed by the oxidation 
of copper, and melts at a red heat without decomposition ; further 
lieating in the presence of air produces the cupric oxide which is less 
fusible. As has been already indicated, cuprous oxide dissolves in 
the molten metal. It is easily reduced to metallic copper by lieating 
w ith carbon, the metal being also obtained if the oxide be heated in 
the presence of reducing gases ; it combines readily with silica when 
lieated, yielding fusible silicates. 

When cuprous oxide is heated with sulphide of iron, the copper, 
having a greater afiinity for the sulphur than iron possesses, enters 
into combination with it, forming copper sulphide and iron oxide, 
and if sufficient silica be present, a silicate of iron slag is produced. 
When melted with copper sulphide, cuprous oxide yields metallic 
>pper with liberation of sulphur dioxide according to the equation — 

CU2S + 2CU2O -> 6Cu + SO2. 

This reaction is a quantitative one, and takes place in the Direct 
Process of NichoU and James as operated at Swansea. The excess 
of either constituent remains unchanged. The reaction is of great 
importance in the processes of copper extraction, since upon it 
depends the liberation of metallic copper from the sulphide, both in 
the old roaster process and in the modern converter operation. 

Sulphides of Copper. — Of the sulphides Cu^S and CuS, the former 
'uly is of metallurgical importance. It is grey black, brittle, and 


crystalline, its melting point is about 1,135° C, and its specific 
gravity, when cold, about 5-5. Owing to the great affinity of sulphur 
for copper, this element acts as practically the universal carrier of 
the metal in smelting work, detaching the copper from all other forms 
of combination, and collecting it as sulphide, mixed with the sulphides 
of other metals, particularly that of iron — copper sulphide and iron 
sulphide alloying in all proportions. 

When copper sulphide is melted with an excess of sulphur, it 
remains unchanged ; when melted with copper and subsequently 
cooled, the sulphide and metal separate as such, although it is 
believed that small amounts of copper are present in solid solution in 
the sulphide on solidification, but that they separate from it during a 
dimorphic change in the material, which occurs at about 103° C. 
The sulphide reacts with iron with liberation of some metallic copper, 
and the formation of some iron sulphide which associates itself with 
the rest of the copper sulphide, forming a matte. This matte is 
not further affected by iron, so that it is not possible to completely 
decompose copper sulphide by this means. 

When heated in a powdered condition in excess of air, copper 
sulphide is oxidised, oxides of copper and sulphur being produced. 
There occur probably several intermediate reactions, and several 
intermediate products are formed, but the main effect is represented 
by the equations — 

( Cu,S + 20 -> 2Cu + SO^ 
I 2Cu + O "> Cu,0 

which take place simultaneously, the copper represented in the first 
equation being oxidised spontaneously according to the second, and 
the resultant is the reaction Cu^S + 30 -> Cu^O + SO.,. 

In the furnace operations, some of the SO.2 in the presence of air 
and oxidisable material, and in contact with the heated brickwork 
becomes oxidised to SO3, which, interacting with the oxides and 
sulphides present, combines to form copper sulphate and cupric 
oxide. At a higher temperature the sulphate is again decomposed 
to CuO and SO3, some of which passes off and is free to oxidise more 
sulphide ; the rest is decomposed to SO., and oxygen. These reactions 
occur during the roasting of charges containing copper sulphides. 

Copper Mattes, — On smelting a furnace charge which contains 
both copper and sulphur, the sulphur appears to have a stronger 
attraction for the copper than for any of the other metals usually 
present, and only when this affinity has been satisfied is the excess 
.sulphur free to combine with other constituents of the charge. The 



fusible copper sulpliide which is thus produced, hcas the power of 
mixing completely with any more sulphides which may be present, 
especially with sulphide of iron. 

The fused sulphides resulting from such furnace operations are 
termed copper-mattes. The^ may contain from a mere trace to up- 
wards of 80 per cent, of copper, and in ordinary work, sulphide of iron 
/ is the other constituent present in the greatest proportion, but sul- 
phides of nickel, silver, zinc, or lead, etc., may also be found, as well as 
arsenides and antimonides. 

^These facts relating to the collection of the copper as a constituent 
of a fused sulphide product, form the basis of modern copper-smelting 
^^ork. In view of the practical importance of the mixed sulphides, 
the diagram representing their equilibrium requires notice. A number 
of workers have studied the question with Avidely dijffering results. 
Rontgen made an exhaustive investigation of the system FeS— Cu.^S, 
and pubhshed a very complete diagram of the series, working Avith 
Fe8 and Cu^S in the pure state. 

The sulphides as commonly met with, especially in smelting 
practice, do not however, occur as materials of the composition 
denoted by the formulae Cu.^S and FeS. The ordinary commercial 
sulphide of iron corresponds more closely to the impure eutectic of 
the iron-FeS system, containing about 85 per cent, of FeS and 15 per 
cent, of iron, and melts at about 970° C, whereas the pure FeS has a 
melting point of upwards of 1,180° C. At the elevated temperatures 
of the copper-smelting furnace, pure FeS tends to lose sulphur and 
to assume the composition of the eutectic. There are, further, good 
reasons for believing that copper sulphide behaves in a somewhat 
similar manner, so that the series of sulphides constituting the mattes 
of practice are not represented by pure materials so well as by a series 
composed of mixtures of the respective eutectics. 

The diagram of this series of industrial sulphides was worked out 
by Hofman, Caypless, and Harrington, and gives a fair summary of 
the melting points of the series of mattes. It is reproduced in fig. S. 
The temperatures may be supplemented by Gibb's determinations of 
1,121° C. for the 71-7 per cent, copper matte, and 1,098° C. for the 
BO per cent, matte. 

The problem of the constitution of mattes is, however, a very 
'mplex one. and is not yet satisfactorily settled. An interesting 
lew was put forward by Gibb and Philp. Mattes corresponding to the 
formula 5Cu.^S . FeS (copper 71-7 per cent.), when examined micro- 
scopically, appeared to be homogeneous, and indicated some form of 
'mbination between the sulphides in these proportions. Lower-grade 



mattes were assumed to consist of this compound substance and 
excess FeS. Iron sulphide was held to be capable of carrying a certain 
(juantit}^ of copper in solution, and mattes might, therefore, carry this 
copper, according to the amount of excess FeS which they contained. 
Within certain limits the lower the grade of the matte — i.e., the more 
FeS present — the more copper was held in solution, and with a fall of 
temperature this sohibility was lessened, and moss copper Avas set free 
in the solid matte. 

Deposition of copper may also be accounted for by a variation in 
the solubility for copper, accompanying the well-marked dimorphic 
change occurring in FeS at 130° C. whilst another possible cause of 













§ ^^^ 




























Per Cent FeS 










BOPer Cent Copper 

Composition of Matte. 

Fig. 8.— Freezing-Point Curve of Iron-Copper Sulphides (Mattes). 

the separation of moss copper is the partial decomposition of Cu^S, \ 
being effected, as previously indicated, by the free iron of the iron-FeS 
eutectic which constitutes the iron sulphide component of copper 
mattes. The whole subject is thus of considerable complexity, and 
involves questions of thermal and chemical equihbrium. 

The appearance, chemical constitution, and physical properties of 
mattes vary according to the rate of cooling, and are further influenced 
by the nature and amount of the impurities they contain, and the 
following statement must be understood to be more or less general ; — 
Usually low-grade mattes (up to 20 per cent, or so of copper) are more 



or less stony in fracture, with a bluish -purple colour ; as the copper 
contents increase, a reddening of the colour occurs, and also an 
increase in the crystalline character and brittleness. Considerable 
quantities of moss copper are present in these mattes. Beyond 30 per 
cent, of copper, increased softness and brittleness result, with a 
darkening towards blue-black in the colour, whilst with the 60 to 70 
\)er cent, mattes the colour becomes in general of a steel-grey hue. 

Increase in the copper contents leads to an increase in the density 
— a matter which has important applications in connection with 
the economical separation of matte from slag, and the slag-losses in 
smelting practice. 

The specific gravity of the 13 per cent, copper matte is about 4-80. 

43 „ „ „ 5- 18. 

60 „ ., „ 5-42. 

80 ,, „ „ 5-55. 

(Gibb and Philp.) 

The density in the fluid state, which is the important condition in 
.^melting, Ls less than this, and may indeed be somewhat different, 
owing to changes in the constitution of the material. 

Copper Silicaie is formed by the action of copper oxide and silica 
II heating. The silicate is decomposed when heated in the presence 
t sulphides, resulting in the formation of sulphide of copper and 
ilicate of the second metal, in consequence of the great affinity of 
opper and sulphur. Upon this fact depends the extraction of copper 
! om various silicate ores, as well as the cleaning of slags high in 
opper. which are often added to the sulphide charges in the furnace 
.. ith this object. When heated with iron, the silicate is reduced to 
metallic copper with the production of silicate of iron ; it is also 
'duced by carbon in the presence of metallic oxides capable of 
ifiiting with the silica which is liberated. 

The Varieties of Commercial Copper, — The copper employed 
industrially comes into the market in widely differing forms. Different 
varieties are named according to the method of manufacture, the uses 
for which they are intended, the locality in wh'ch they are produced, 
or by special trade names. The most imports* xt variety is : — 

Ehctrolytically -refined High-conductivity Copper, which is largely 
used for electrical work. The methods by hich it is produced ensure 
that most of the impurities inimical to high conductivity have been 
removed, and the metal is specially free from arsenic, antimony, and 
bismuth, as well as from silver and ^r\d. As ordinarily produced at 
the electrolytic refinery, it is in tlie form of cathode plates, often 



about 3 feet x 2 feet 6 inches by J inch thick, weighing 150 to 170 
lbs. It is then remelted in order to bring it " up to pitch,"' and to give 
it the necessary mechanical properties, so that it may be transformed 
at once into the particular form suitable for the electrical purposes 
intended. Such metal often comes into the market in the form of 
wire-bar ingots, cakes, or billets, weighing from 70 to 500 lbs. when 
in bar form, and from 100 to 400 lbs. when in other shapes. 
Electrolytic copper is also suitable for the manufacture of alloys. 

Lake Copper. — The copper ores of the Lake Superior district aie 
particularly pure, and on smelting and furnace-refining yield a metallic 
product of great purity which also possesses good mechanical pro- 
perties. Tt is, therefore, particularly suitable for electrical work. By 
reason of its satisfactory properties, Lake copper realises prices Avhich 
usually rule somewhat his/her than those of ordinary electrolytic 
copper as quoted on the Ncav York market. 

Be-H Select Copper. — For the production of copper alloys, sucli as 
best brass, etc., it is essential that the copper should be pure. The 
impurities which are present in ordinary tough copper, and which may 
be valuable for imparting strength to the material, have a very 
harmful effect when present in alloys. In the older Welsh process of 
manufacturing copper, a special method Avas employed for obtain in j^ 
metal free from these impurities, especially arsenic and antimony. 
This was known as the " best selecting " process. 

The principle underlying the method was to conduct the furnace 
operations to the stage at which a small quantity of copper, known 
as '' copper bottoms," was obtained. The metal so produced has the 
property of collecting from the rest of the matte-charge in the furnace, 
the gold, the silver, and the great bulk of the other impurities., owing 
to its greater solvent power for them. As a result, the greater part 
of the matte (''white metal") was left pure, and from this material 
the copper was extracted by continuing the furnace operations in 
the usual manner, the resulting product being known as " best select " 
(B.S.) copper. 

The process was later used principally for the extraction of the 
gold in the charge, rcther than for o})taining specially pure copper. 
The product is essentii^^ly a British one, and was largely used for the 
manufacture of high quf) Hy alloys. 

'*' Tough Pitch Coppet ' — The operation of " bringing copper up 
to pitch " has for its object ' he imparting to the metal of the toughness 
and mechanical strength required for industrial service. The process 
resolves itself into the adjustm. nt of the correct proportion of oxygen, 
the function of which is largely to eliminate the gases from the copper 



or to ovej'come their deleterious effects, as Avell as to convert the 
otherwise more injurious metalloid impurities into a less harmful form. 

In modern practice, practically all copper is brought up to pitch, 
but it is useful to distinguish between tough-pitch furnace-refined 
copper and tough -pitch electrolytic copper. 

The former is the brand to which the general term '' tough pitch 
copper ** is best applied, this name having been given to the product 
from the refinincr furnaces of the old Welsh and similar processes. 
Before the converter method was introduced into copper practice, 
the furnace processes for extracting copper from the ores resulted in 
the production of a crude " blister " copper, into which several injurious 
constituents, if originally present in the ore, found their way. The 
principal impurity was usually arsenic. Although this was also re- 
movable by special refining methods, and with some difficulty, it was 
known, as has been indicated, that when arsenic is present under 
suitable conditions and in proper proportions, it is capable of im- 
parting considerable strength and rigidity to the metal. Such copper 
being particularly suited for various engineering and mechanical uses, 
the arsenic being sometimes even specified for and purposely added 
— as in fire-box plates and stay bolts, though it is never employed for 
conductivity work or for the manufacture of alloys if any considerable 
proportion be present — the metal found a ready market when brought 
to pitch. 

Tough pitch copper may thus vary largely in composition, 
especially in arsenical contents, up to about the 0-5 ])er cent, already 
indicated as being mechanically very useful. The actual process, as 
used for bringing all classes of metal to pitch, will be described in detail 
later, it being practically the same whether conducted on furnace- 
refined metal, converter metal, or on electrolytic copper, as a necessary 
preliminary to casting into the various forms of ingot in which it is 
to be marketed. 

In preparing the tough metal from crude copper, the more oxidis- 
able impurities (iron, sulphur, etc.) are first removed by a thorough 
oxidation during or after melting down, this being known as " airing.'' 
The operation oxidises some of the copper, and it is probable that the 
copper oxide thus formed plays an important part in getting rid of 
impurities. By the time they have been thoroughly expelled, the 
metal is considerably over-oxidised. Samples taken at this stage 
exhibit the following characteristics : — The ingot has a depression 
down the centre line, the material is very brittle, the fracture is brick- 
like in texture and purple-red in colour, whilst much copper oxide and 
xidule-eutectic are seen on examination under the microscope. This 


material is known as Dry Co27per ; it is merely an intermediate product, 
and is commercially useless. The excess of oxygen is removed by 
''poling'' — that is, reduction, effected largely by charcoal, as well as 
by reducing gases — successive samples showing less and less of the 
characteristics of dry copper. The surface becomes level, the metal 
exceedingly tough, the fracture fine-grained to silky in texture, 
and a fine salmon -pink in colour. With satisfactory mechanical 
properties, the metal has now become tough pitch copper. 

Tf the poling — that is, the reduction of the oxidised constituents 
of the tough pitch copper — be carried too far, the metal becomes 
brittle again, beins^ known as over-poled copper. The fracture then 
tends to become coarse and fibrous, the colour lighter, and the upper 
surface of the ingot exhibits a ridge. The reasons for these effects 
have not yet been quite fully explained, but there is no doubt that 
the}^ arise from the removal of oxygen from the oxygenated con- 
stituents, and the withdrawal from the metal of the protecting in- 
fluence of the cuprous oxide. Such influences are to some extent 
physical, since they prevent the retention of the reducing gases; 
partly mechanical, in their effects on the properties of the metal per se, 
and partly chemical, as the oxide had probably entered into chemical 
combination with some of the objectionable impurities, producing com- 
pounds, in which form they were much less harmful. The removal of 
this oxygen from the metal breaks down such combinations, leaving 
the reduced impurity again to exercise its destructive effects on the 
properties. Over-poled copper, like dry copper, being brittle, is com- 
mercially useless as such, and is really an intermediate product, the 
metal being brought to pitch again by further aeration to make it 
" dry,'* after which it may be poled back to correct pitch. As already 
stated, the over-poling effects are not due to any intrinsic action of 
carbon directly on the copper itself. 

Summarising, it may be stated that the most important com- 
mercial varieties of copper are : — 

Electpolytically-refined metal, employed for electrical work (also 
for alloy-making). 

Tough Pitch Copper for engineering uses. 

Best Select Copper for alloy manufacture. 

And, in addition. Lake Copper and some Converter Bars. 

A number of unrefined metallic products met with in practice 
include : — 

Converter Bars. — The product from the Bessemer operation on 
copper mattes. Most converter metal is subsequently electrolytically 
refined, but several varieties of Australian and American copper are 




]nit on the market direct in this form. Being produced from fairly 
pure ores, which carry but Httle silver and gold, the converter metal 
may be sufficiently pure to render electrolytic refining unnecessar3^ and 
too low in gold and silver values to make such an operation profitable. 

Cathode Copper is the product from the electrolytic refinery, and 
is usually remelted. brought up to pitch, and cast into ingots previous 
to use. 

Blade Copper is produced by the smelting of oxide ores, and is 
subsequently refined. 

Cement Copper is produced by wet processes, usually by precipi- 
tation from copper-bearing solutions bj^ means of iron, the product 
/being a rather impure reddish-brown spongy mass. Many varieties 
y contaiaarsenic. It requires melting and subsequent refining to adapt 
Ttlor service. 

Blister Copper w as the name given to the crude metal from the 
older type of furnace operations. Such copper contained large quanti- 
ties of gas, particularly S0.>, which, tending to escape at the moment 
of solidification in the mould, gave a bUstered appearance to the 
surface. It contained 96 to 98 per cent, of metallic copper, and was 
subsequently refined. The term is generally applied still to all crude 
' opper exhibiting similar features. 

Chili Bar is an impure copper imported from Chili for refining. 
I he composition varies, the metal usually containing 96 to 98 per cent. 
f copper, A^ith indefinite quantities, sometimes small, of undesirable 

Appended is a series of representative analyses of various copper 
products, compiled from different sources. The composition of such 
material as tough pitch copper and the various cruder varieties is, 
however, subject to very great variation. 

The Sources of Copper.— Copper ores usually consist of various 
minerals of copper mixed with those of many other metals, and accom- 
panied by very varied gangue, according to the locality in which they 
re found. 

They are best classified under three groups : — 

(1) Native Ores. 

(2) Sulphide Ores. 

(3) Osiide Ores, 

The most important points to be noted with regard to the dis- 
tribution of these different classes are that — 

(1) Native ores are localised in their occurrence, being chiefly 
'mfined to the Lake Superior district. 





(2) Sulphide ores supply the bulk of the world's copper, constituting 
upA^ards of 80 per cent, of the total. 

(3) The oxidised ores are found in most copper districts, thougli 
usually to only a limited extent. They are often gossan deposits 
produced by weathering or by decomposition of sulphides, hence 
are generally found nearer the surface, changing to sulphide with 
depth. The supply of copper from oxidised ores, which was at one 
time very large, is decreasing rapidty, and the greater proportion of 
the copper now obtained from them comes from the more recently 
developed deposits, of which those at Tanganyika afford an example. 

]More than 200 minerals which contain copper are known, but 
most of them are unimportant from the smelting point of view. The 
characteristics of the more noteworthy may be fully studied from 
text-books of economic mineralogy. 

Copper Ores — Native Copper, — Occurs extensively in the Lake 
Superior district of Michigan, in Precambrian rocks, sparingly in New 
Mexico and China, but seldom anywhere else in workable quantities 
by itself. Copper barilla or copper sand, an impure native metal from 
Chili, was formerly of importance. Native copper constitutes about 
20 per cent, of the North American supply. It yields metal of ex- 
ceptional purity, and the brands of Lake copper reach a very high 
standard, both as regards electrical and mechanical properties. 
A still purer variety is the native metal from Yunnan, China. 

The Lake Superior copper occurs in three formations : — 

(a) Vein deposits, from which the enormous masses of copper are 
taken out. 

(h) Copper-bearing ash beds, of amygdaloidal diabase. Chief mine, 

(c) Beds of conglomerate in which the cementing material consists 
partly of copper. This last class of deposits yields three-quarters of 
the Lake copper supply. Their average copper content is 2-9 per 
cent. The chief mines are the Calumet and Hecla, the Tamarack 
and the Atlantic, all situated on one ore chute measuring 3 miles in 
length, and worked to a depth of 4,000 feet. 

Sulphide Ores : Chalcopyrite (Copper Pyrites) is by far the most 
widely distributed ore of copper, and furnishes the greater proportion 
of the world's supply. 

The formula when pure is Cu,S . Fe,S, (Cu 34-4, Fe 30-6, and 8 
35-1 per cent ), but usually the ore is not in this condition, being 
mechanically mixed with large quantities of iron pyrites, and very 
often with pyrhottite. It occurs principally in the older crystaUine 
Tockn, often in bedded veins. 


The value of copper veins below the limit of surface decomposition 
is nearly always due to chalcopyrite. Silver and gold are often carried, 
as well as other metals. It occurs extensively in Montana, Arizona, 
Tennessee, Canada, Chili, Japan, Spain, Cornwall, etc. 

Chalcopyrite ores vary considerably in copper contents : thus 
Tennessee ores contain about 2-5 per cent, of copper, Montana ores 
o to 5J per cent, (with gold and silver valued at about £11 per ton of 
copper), whilst the Arizona ores vary, being often rich. 

Chalcocile (also known as copper glance or redruthite) is much 
less important. The copper contents are 79- S per cent, when pure, 
but such a condition is rare, although the ore seldom contains less 
than 50 per cent, of copper. Below this proportion it often tends to 
pass into bornite, and then to chalcopyrite, It is found in Montana, 
is an important ore in Arizona (Clifton district), and occurs also in 
Cornwall . 

Other important sulphides include : — 
- Borniie (Erubescite, Peacock copper ore), 3Cu._,S . Fe.^8... occurring 
in Cornwall, which passes with depth into chalcopyrite 

Tetra?iedrite (Fahl ore), a very complex sulphide of copper, iron, 
lead, zinc, with arsenic, etc. It is often rich, and carries silver values. 

Oxidised Ores. — The most important of the oxidised ores are — 

Malachite, CUCO3 . Cu(OH)o, containing, when pure, 57-3 per cent, 
copper (73-7 CuO) ; is widely distributed, but usually occurs as such 
in non-paying quantities except in a few particular localities. It is 
found in the upper parts of the veins. Whenever found with sul])hide 
ores, it is an extremely useful material to mix in the charge, as i 
supplies oxygen as well as copper. Malachite is still an important 
source of the metal in Mexico, ChiU, and Bolivia, though not quite 
so much so as formerly, whilst it is specially important in the Tan- 
ganyika (Katanga) deposits, of which it constitutes the greater portion 
so far developed. 

Cuprite, Cu^O, contains 88-8 per cent, copper, when pure. It is 
Avidely distributed, but is never found by itself in paying deposits, 
though in the early days of mining and smelting it was an important 
source of metal, since it was easily reduced, and consequently A\as 
cheaply worked. 

Melaconite, CuO, contains 79-8 per cent, copper, when pure ; is fairly 
widely distributed, although hardly ever in sufficient quantity to pay. 
In one or two localities, however — viz., Tennessee, North Carolina, 
and Virginia — it was formerly an important source of the metal. 
The deposits were at first very promising, as they consisted largely of 
very rich melaconite ore ; this was however, quickly worked out, the 




ordinary heavy chalcopyrite with 2-5 per cent, copper being struck 

Other oxidised ores include — 

Azi'rite. 2CUCO3 . Cu(OH),, and Atucamite, CuCl., . 3Cu(0H).,. from 

In modern work, the chief ore smelted is impure chalcopyrite. 
Carbonate and oxidised ores, when they can be obtained, are mixed 
with it, increasing the concentration and shortening the process ; 
except under certain special circumstances. 

Preliminary Treatment of Ores. — The treatment of ores pre- 
paratory to smelting includes the processes of sampling, wet concen- 
tration, agglomeration of fines, and roasting. 

Sampling.— Since sampling is not a part of the extraction process 
proper, in copper smelting, it will be convenient to deal Avith the 
subject separately here. 

It is important that ores and all other products entering or 
leaving the works, as well as many of the intermediate products of the 
various operations, should be properly sampled and assayed. Great 
attention is paid to this point at the best organised smelters, since 
only by this means can the work of the plant and of its several depart- 
ments be properly checked and controlled. Each works has its own 
special method of taking samples from the stocks, the Anaconda 
practice, for example, being to pass the whole of the first-class ore, 
amounting in quantity to 25,000 tons per month, through the samp- 
ling mill, whilst of the poorer, second-class, ore for concentrating, 
every fifth car-load is sampled. 

There are many different types of sampling plant, and the 
methods employed vary also, but the principle is much the same in 
each case — namely, to use some automatic device which cuts out 
and deflects a certain proportion of the stream of ore on its course 
through the mill ; — the deflected portion being crushed finer, and a 
part of it again cut out and deflected ; repeating the operation in this 
way three or four times. 

The sampling process and plant at Anaconda is so representative 
t the best practice, that it may be reviewed in brief, as an example. 

The Anaconda Sampling Plant is entirely automatic in its action. 
The mill is built in two sections, each of which treats 1,800 tons daily. 
Kach section consists of a set of four sampling machines with inter- 
mediate crushers. The ore goes from the bin to a Blake crusher, 
breaking to 3-inch to 4-inch size ; the crushed ore is elevated and fed 
down a chute to the first sample cutter, which takes out one-fifth 


a — 1 




^^20(11 by 10 in 

y 1st Cut -too iniOCJlb 

/^ I \ n Crusher 
L-UJ (15" X 5) 

\ ?nd Cut^iii-IOOIb 

/ C^V^ Fin*- Ro'ls 

Srd Cut !6in80 11> 

(y^ Sample Rolls 

4th Cut 3-2 111 16 lln 

/ Sample tofinisli 


a — 

j^ia. 9. — Outline of Sampling Scheme, Anaconda. 

CouDt«r Shaft 

=«00 LB. SAMPLE 

So. I Sampler 

16 X 3C RoiU 
Nu. 2 Sampkr 
2N0 CUT 
-ao LB. SAMPU 
4 X 27 RoU» 
= 16 LB. SAMPLE 

X X Rolls 
Line Shaft 

= 3.2-LB. SAMPLE 
\j- Sample Safe 

Fig. 10. — Section through Sampling Mill. 



(400 lbs. per ton) as a sample, and deflects the rest down another 
chute. The sample is crushed further in a Blake crusher, and passes 
a second sample cutter (rather smaller in size), which again takes 
out one-fifth (80 lbs. for every original ton of ore), and rejects the 
rest down the " rejects-chute." The sample is now crushed in rolls, 
a third cut of one-fifth (16 lbs. of the ton) taken as before, the rest 
rejected. The sample passes to a final set of crushing rolls, and the 
last cut of one-fifth is taken. Hence each ton of ore is represented 
eventually by 3-2 lbs. of sample. 

The sample cutter employed is of the Brunton form. It consists 
essentially of a curved boat of 120° arc, which rotates to and fro on 
a central spindle. The top is open ; one side has one hole cut in, the 
other has two, the area of the latter being together four times that of 
the single one, so that the falling stream is cut continually, and -quo 

Fig. 11. — Brunton Sampler. 

fifth is deflected to one side, falling down a chute to the next crusher, 
whilst the other four-fifths fall from the other side to the rejects- 

The above description is quite general, several details for certain 
classes of ore having been omitted, but it gives a fair idea of the 
general principles underlying such work. 

The final sample, say 3,200 lbs. per 1,000 tons of ore, is mixed 
on an iron plate on the floor, quartered several times by a Brunton 
shovel, and the chosen sample then ground in an Englehardt mill 
(small Gates' crusher with two discharges). The material is passed 
^irough a 1-foot riddle of 100 mesh wire cloth, the very small quantity 
i coarser stuff remaining, is bucked down and added, and the 
hole is then thoroughly mixed in a canister of 1 foot side gripped 

opposite comers, and rotated mechanically. 




Constitution of Copper Mattes. 

Keller. Mineral Industry, vol. ix., 1900, p 240. "Elimination of Impurities from 

Copper Mattes." 
Rontgen, Metallurgie, vol. iii., 1906, p. 479. 
Hofman, Caypless, and Harrington. Trans. Amer. Inst. Min. En/j., vol. xxxviii., 1908, 

pp. 142-153. 
Gibb and Philp. Trans. Amer. Inst. Min. Eng., vol. xxxvi., 1906, p. 665. 
He37n and Bauer. Metallurgie, vol. iii., 1906, p. 84. 

Fulton and Goodner. Trans. Amer. Inst. Min. En^., vol. xxxix., 1908, pp. 584-020. 
Refining of Copper. 

H. 0. Hofman, R. Hayden, and H. B. Hallowell, " A Study iu the Refining and Over- 
poling of Electrolytic Copper. Trans. Inst. Amer. Min. Eng. 
Hofman, Green, and Yerxa, " A Laboratory Study of the Stages in Refining Copper. 

Trans. Amer. Inst. Min. Eng., 1904, vol. xxxiv., pp. 671-95. 

Stahl, " Ueber Raffination, and Eigenschaften des Kupfers." Berg, and HutteU' 

m.anmsche Zeitung, 1889, vol. xlviii., pp. 323-4; 1890, vol. xlix., p. 399; 1893,^ 

vol. Iii., p. 19 ; 1901, vol. Ix., pp. 77-79. 

Keller. Mineral Industry, vol. vii., p. 245, et seq. 


D. W. Brunton, " Modern Practice in Ore Sampling." Mining and Scient. Press, Oct. 30, 

" Theory and Practice of Ore Sampling." Trans. Amer. Inst. Min. Eng., vol. 
XXV., p. 826. 



lYIodcrn Copper Smelting- Practice— Preliminary Treat- 
ment of Ores: Concentration, Briquetting-, Sintering* 
—The Principles of Copper Smelting— Roasting. 

Modern Copper Smelting Practice* — Until recently, modern 
smelting practice has been understood to involve the production of 
a matt^ containing from 40 to 50 per cent, of copper, which is then 

There are however proceeding at present (owing to the successful 
working of basic-lined converters) developments which indicate that 
-uch practice may. within a few years, be modified very considerably 
in the direction of the converter treatment of lower-grade mattes. 
Until such operations become successfully established and generally 
adopted, the production and subsequent bessemerising of 40 to 50 per 
cent, matte will be here dealt with as constituting modern practice ; 
particularly since, generally speaking, the principles involved are 
equall}^ applicable to the modified methods now being developed. 

Preliminary Treatment of the Ore, — The factors which have 
to be considered in drawing up a scheme of treatment for the supply 
of ores shipped to a smelter are exceedingly numerous, and will be 
discussed in due order. There are no hard and fast principles which 
determine such schemes, yet a number of considerations must be 
noted concerning the treatment preliminary to the actual smelting 
of the ores. 

Such preliminary treatment may include — 

A. Concentration or Wet Dressing, 

B. Agglomeration of Fines — (a) Briquetting, (b) Sintering. 

C. Roasting. 

A. Concentration or Wet Dressing. — In treating the ores of copper, 
it may be noted that in general — 

Native Ores, unless very massive, are usually dressed in a special 
manner peculiar to themselves — e.g., stamp-miUing. 

Oxide Ores are rarely wet-dressed. They present much difficulty 
n treatment on account of their comparatively low density, which 
iiakes efficient wet concentration almost impossible, whilst heavy 
losses in the tailings generally accompany such operations. 



Sulphide Ores. — No definite rules can be laid down as to whether 
the ore should be wet-dressed or not ; the treatment depends 
altogether on attendant circumstances, such as — (a) the character 
of the ore, (h) the concentration of the copper desired in the first 
smelting operation, and (c) the smelting method and furnaces 

Wet concentration is only profitable Avhen the copper ore is of low 
grade, and then only under suitable conditions. Thus the low tenor 
may be due to admixture with much gangue or with other sulphides, 
or both. A massive low-grade pyritic ore carrying but little gangue 
is not suitable for such treatment, since the mixed sulphides are not 
separated from one another by wet dressing, and consequently but little 
enrichment of the copper in the dressed product would be possible ; 
apart altogether from other considerations. Such is the case, for 
instance, with the Tennessee ores carrying about 2-0 per cent, of copper 
and only 25 to 35 per cent, of gangue. 

An ore with a self-fluxing or almost self-fluxing gangue might 
allow of its copper being concentrated more cheaply and conveniently 
by direct smelting than by wet dressing, this depending, of course, 
on the local conditions. 

In other cases a balance has to be struck as to whether the cir- 
cumstances are more favourable for removing the excess of gangue 
by means of crushing and treatment in a stream of water, or by slagging 
it off in a furnace with the addition of suitable fluxes. In many cases, 
with low-grade ores, the former treatment is the cheaper. 

The case of the low-grade ores of the Butte, Montana, district, 
affords a good example of these considerations. This ore contains 
5 to 5i per cent, copper, with a large quantity of highly siliceous 
gangue. It was found that the purchase and carriage of sufficient 
flux, and the cost of carrying out this fluxing operation was so ex- 
pensive that it was cheaper to build a concentrator and smelter at 
Anaconda, 30 miles away — in a locality where a suitable water supply 
was available for the dressing — and to convey the ore this distance in 
order to concentrate it by a wet method. The dressed ore assays 
9 to 10 per cent, of copper. 

It is important to note that the process of wet dressing involves 
crushing the ore, and yields the product in a more or less finely divided 
form. Most copper sulphide minerals are exceedingly brittle, and 
break up to a very small size on crushing for concentration, so that the 
copper concentrates usually include a large quantity of fine material. 

There are two general types of furnace available for smelting — 
reverberatory furnaces and blast furnaces — and the questions of the 




desirability and of the degree of crushing and concentration depend 
to a large extent on the plant and furnaces adopted or proposed. 

Blast-furnace treatment has hitherto often been considered the 
most economical process for smelting copper ores, especially with 
regard to fuel costs, but for many reasons it is not a convenient or 
efficient furnace for the direct treatment of fine material. When it 
is desired to employ the blast furnace, it is necessary to make up 
charges consisting, to as great an extent as possible, of coarse material. 
In consequence, when concentrating ores with a view to subsequent 
blast-furnace treatment, the degree of crushing and dressing has 
to be modified with these factors in view ; otherwise a further preH- 
minary manipulation of the fine concentrates that are produced is 
rendered necessary. Such modified dressing schemes involve a maxi- 
mum of coarse breaking and screening, the crushing and separating 
-tages being thus very gradual, and the units in the plant are 
multiplied, whilst the process is rendered complex in consequence. 
With the greatest care, moreover, large quantities of fines are bound 
to be produced, and have to be dealt with by some means other than 
immediate blast-furnace treatment. 

Dressing schemes and plant for sulphide copper ores are thus often 
complicated, particularly for the recovery of the values from the finer 
material, and cannot be discussed at any length here. Reference 
should be made to Richards or other standard works on the subject. 

As representative of wet-dressing practice, the Anaconda scheme 
may be noted, as summarised below. 

There are eight mills, each treating 1 ,000 tons of ore per day, and 
conducting the — 

Coarse crushing in Blake crushers. 

Coarse sizing by trommels. 

Coarse separation on Harz jigs of l|-inch and |-inch feed. 

Middlings crushing in rolls. 

Middhngs sizing by trommels. 

Middhngs separation on fine jigs of 7, 5, 2J, 1|, and 1 millimetre feed. 

Finest crushing in Huntingdon mills. 

Fines settling by spigot settlers. 

Fines separation on Wilfley tables (471 are in use on the plant). 

The muddy water goes to enormous settling ponds, where the 
slime settles down, gradually drains, and dries, and it is afterwards 
used for various purposes during the smelting operations ; being dug 
out m the form of a fine clay. A new form of centrifugal apparatus 
(the Peck) is now being installed for the separation of this material. 



The subsequent treatment of the products from the concentrating 
operation is indicated in the diagram (fig. 12), from which it will be 
seen that the — 

Coarse Concentrates, IJ, J (and f) inch size, are smelted in the blast 


Concentrating Orej-^^ \ ^ Smelting Ore 



Tail in 





'Calc ines" X^Sli^il 


Slime Ponds 


Briquette Plant 

Briquettes /^ 

gs Slag Sla^ 











Anode C opper 

( Shipped to Eastern Refinerie s) 
Fig. 12. — Outline of Smelting Scheme at the Anaconda Smelter, Montana, U.S.A. 



Fine Concenfrafes, 7, 5, 2 J, IJ, and 1 millimetre size, pass to the 
roasters, and thence to the reverberatory furnaces. 

Slimes are used for briquetting, and several other operations. 

Tailings pass to the dump. 

B. Agglomeration of Fines. — It has just been seen that the wet 
concentration of ores (considered advisable in a large number of 
cases) results in the production of a considerable quantitj'- of fine 
concentrate, a form of material not well suited for immediate blast- 
furnace treatment. 

In addition, smelters often receive considerable amounts of fines 
in the smelting-ore supply, which it is not unusual to screen out and 
to treat separately from the coarser materials.* 

The alternatives for the treatment of fines, and more particularly 
of fine concentrate, include smelting in reverberatory furnaces (usually 
after roasting) ; blowing into the converter (a new process still in the 
experimental stage) ; and blast-furnace treatment after suitable 

Blast furnaces have many advantages which lead to their extended 
use in copper smelting practice, but one important feature, which also 
apphes to the smelting of other metals, has always to be borne in 
mind in this connection — viz., that material in a finely divided state 
cannot he treated directly in a blast furnace without heavy losses, and the 
working of the furnace on such charges is not efficient. 

No material less than J to | inch in size, especially when in the form 
of sulphides, should be fed as such into a modern blast furnace. Fines 
in the furnace lead to — 

(a) Accretions, 

(h) Irregular working of the furnace and' the charge, 

(c) Clotting and loiv concentration, 

(d) Heavy flue-dust losses, 

and their presence is often the cause of much trouble at many of the 
modern smelters. The agglomerating of the fines is, therefore, a very 
important prefiminary in any scheme of treatment involving the 
employment of the blast furnace on such material. Agglomerating 
is usually performed by one of two methods — (I) briquetting, 
^2) sintering. Of these, briquetting has hitherto been in very general 
ise, but several advantages connected with the sintering process and 
• he resulting product are leading to its adoption with much success 
m several localities, and attracting for it considerable attention at 

* The treatment of flue-dust is considered later. 



(a) Briquetting, — Among the advantages of briquetting is the fact 
that it utilises large quantities of the copper-bearing slime produced at 

Briquette Bins 

Con veyoF 




WM. ^^^' 




FeedChut^ Sl/m^S 

Ore Screening 


- 1 Chut 

Chute — 


Con v^ or 



^Charging Platform 

Belt Conveyor 

Fig. 13.— Sketch Plan of Briquetting Plant. 

Fig. 14. — Section through Auger- Former, showing Briquetting Mechanism, of 
Chambers' Machine. 

the concentrating plant, this material often possessing good binding 
properties which render it very suitable for briquette-making. 




The type of plant in use at different smelters varies considerably, 
the method adopted being either the stamping out of the briquettes, or 
by the application of steady pressure, the production of bars which 
are then cut up to convenient size. 

The constituents used depend naturally on the materials available 
at the smelter, briquettes, both with lime and without, being made. 

The Briquetting Plant at Anaconda. — The operation of this plant 
affords a good example of the process. Its working is very successful 
in using up much fine concentrate, as well as the slime from the ponds, 
which acts as binding material and at the same time supplies copper. 
Briquette, indeed, constitutes one of the biggest items of the charge 
for the Anaconda blast furnaces. There are four Chambers' machines 
in use, making 840 tons of briquettes daily. The briquettes consist 
of slime, fine first-class ore screenings « f-inch size), fine concentrate 
from the dressing plant, and coke (which is recovered from the 
reverberatory furnace gratings). The quantities used daily are some- 
what as follows, though they are naturally subject to some variation, 
depending on supplies : — 

Slime, ..... 


. 500 tons. 

First-class ore screenings, 


. 300 „ 

Fine concentrate, . 


. 200 „ 

Coke, , 


. 70 „ 

and the composition of the briquettes is 



Copper, .... 

5-0 per cent. 

Ferrous Oxide, 


Silica, ..... 

45 to 50 „ 

Sulphur, .... 


Lime, ..... 


Moisture, .... 




The different materials are stored in bins, and fed through doors to 
conveyors, which discharge on to an elevator leading to a divided 
hopper, each division of which feeds a pug-mill. The pug mills are 
long troughs in which inter-moving bladed spindles rotate, churning 
tip the materials ; the mixing being assisted by a water supply from 
above. The mixture passes down a chute to one end of an auger 
machine, from which it issues, through a steel ring, in the form of a 
continuous slab, 6 inches X 4 inches in section, to a cutter 10 feet 
distant, which slices off bricks 10 inches long, each of which weighs 
about 10 lbs. The bricks pass to a traveller, thence by another to 


feed bins. The briquettes are not dried, but are used just as made 
with 15 per cent, of moisture, and are generally the last item of the 
charge to be added on the car. They crumble slightly, but are 
sufficiently strong to stand the handling during charging. 

Many similar methods, including hand processes, are emploj^ed. 

(b) Sintering Processes. — This method of treating fines involves roasting 
reactions, as well as the mechanical process of agglomerating. Whilst 
it thus furthers the concentration obtained in the subsequent furnace 
operation, since it eliminates some sulphur, it also utilises the fuel 
value of the fines, and yields a product which works well in 
the blast furnace. Several processes have been introduced, and the 
M'Murty-Rogers method installed at Wallaroo, S. Australia, illustrates 
very well the principles upon which this class of treatment depends. 
It is a sintering and roasting process similar in type to the Huntingdon- 
Heberlein method for lead smelting, but lime is not used as a rule. 
It is employed primarily for fine concentrates which are somewhat 

Charge. — Must contain 15 to 35 per cent, sihca, and 15 to 25 per 
cent, sulphur. 

Pots. — 8 feet 6 inches in diameter, when used for ore, and 4 feet 
6 inches deep ; with vertical sides. There is a false grate 10 inches 
above the bottom, pierced with f-inch holes. 

Blast. — 1,000 cubic feet per minute at 13 to 20 ozs. pressure per 
square inch. 

Capacity, 8 to 10 tons. Time, 8 to 10 hours. 

Method. — Cover the grate with a layer of roasted material, light 
small fire of wood, blow, and gradually charge in the ore whilst the 
blast is on. Lime is unnecessary, but water is essential in the process, 
and the ore must be very wet ; 6 to 9 per cent, water being used for 
ore charges, and 3 to 4 per cent, with rich mattes, otherwise working 
is not uniform, and the losses by dusting are great. With the requisite 
quantity of water present, the working is regular and uniform, there is 
little dust, and the roasting is efficiently performed. 

Products. — If ore is charged, a sintered mass of matte and ferrous 
silicate results ; if poor matte is used, the product is a rich matte and 
ferrous silicate ; and if rich matte is used, metallic copper and ferrous 
silicate are obtained. At the end of the blow the charge is tipped out 
and fed into the blast furnace. 

Costs. — The method as employed at Wallaroo to treat 400 to 500 
tons of material per week, operated at a cost of 3s. 6d. per ton, or about ' 
Is. more per ton than for ordinary roasting. 



Though this particular process is only, to the author's knowledge, 
employed at a fcAv smelters, sintering or blast-roasting methods on the 
same principle have been introduced at several other works, and 
their adoption promises to lead to very successful results, being par- 
ticularly suited for the class of material indicated above. The advan- 
tages claimed for the process are that — 

(a) It saves heavy mechanical losses, such as those of the dust 
resulting from calcining operations and from the charging of hot 
calcines into reverberator}^ furnaces. 

(/>) It gives a product suitable for blast-furnace smelting — often 
the cheapest and most convenient method of working. 

(r) It results in efficient roasting and good reduction of sulphur, 
yields the product in an advantageous form for subsequent smelting, 
and promotes a satisfactory' removal of impurities in the slag. 

In addition, the process offers the possibility in the future of being 
so modified as to leave in the adequately compacted products so much 
sulphide that their fuel values can be reahsed in the blast furnace 
In other words, after the preliminary sintering process, to smelt the 
(fine) sulphide-concentrates pyritically in the blast furnace. 

Of the more recent types of machine for conducting the process of 
.sintering, that of Dwight and Lloyd is in operation at several 
smelters. The moistened ore falls on to an endless chain conveyor, 
composed of separate grids carried on wheels. The conveyor carries 
the ore through the flame from a small furnace which starts its 
ignition, and it is then drawn over a long suction chamber where air is 
sucked through the hot mass, thus effectually roasting and sintering it. 
The chamber has special devices which ensure the drawing in of the air 
through the charge only, and so prevent inward leakage (see Fig. 16). 

The sintered cakes are finally discharged automatically into cars. 
Details regarding the machine vary at different smelters ; at one works 
the length is 30 feet, the rate of travel 8 inches per minute, and the 
vacuum in the suction chamber 6 ozs. 

The size of the particles should not exceed J inch, and not more 
man 25 per cent, of the charge should be so large. Some 3 to 5 per 
cent, tends to pass through the grids, and so be drawn into the suction 
chamber ; this is cleared out at intervals through special doors. 
Water is necessary, and from 6 to 10 per cent, must be employed in 
uniformly moistening the charge, which, by the addition of suitable 
fluxes, is often made of such proportions that in subsequent blast- 
furnace smelting a satisfactory slag is produced without further 
additions. The sulphur reduction by the process is very considerable. 

Such blast-roasting methods, with suitable modifications, promise 








to assume considerable importance in the developments of modern 
smelting practice. 

C. Roasting. — Roasting is often a ver}' important preliminary 
stage in the scheme of treatment of copper ores. It was formerly con- 
sidered an essential operation in smelting processes for sulphide ores, 
the material being crushed and concentrated largely with a view to such 
subsequent treatment. This is not the practice in modern smelting. 
Roasting is now only conducted where the necessity for it arises, as 
in the case where wet dressing, having been considered advisable, 
has resulted in the production of large amounts of fine concentrate, 
and where reverberatory furnaces are installed for the smelting of this 
material. Preliminary roasting of the concentrates then conduces 
to the production of a matte of converter grade in one smelting 

The Principles of Copper Smelting. — Copper extraction from 
sulphide ores is essentially an oxidation process, the iron and sulphur 
being oxidised and the oxide of iron slagged away. All such smelting 
processes, both the older and the more modern ones, are based on this 
fact, and underlying all of them are certain fundamental principles 
which it is essential to keep in mind in considering every phase of the 

These may be summarised as follows : — 

(1) In the melting down of a furnace charge, the copper has first 
claim on any sulphur which may be present. 

(2) Only such sulphur as remains in excess after the copper has 
been satisfied, is free to combine with other constituents of the charge. 

These fundamental principles can best be illustrated by following 
the reactions during the smelting of a typical charge. Thus — 

Copper, • • • Copper. 

Iron, - • •{{ron^xj/' ^^"«^\^ ^^CujS.y.FeS.-A/a^e. 

Sulphur, • 

Oxygen, • • - Oxygen--''' "^ ^^:5 -^xYtO-y^iO^.-Slag. 

t Silica, . • Silicar ->-SiOo''' 

The copper takes up sufficient sulphur to form Cu^S ; the remaining 
Iphur combines with any iron which is available, forming FeS. 
lese two sulphides, dissolving in all proportions, constitute the matte 



The iron in excess of that required by the sulphur becomes oxidised, 
and the resulting oxide combines with silica in the charge, forming 
the silicate slag of the smelting operation* 

It will thus be apparent that, in general, the larger the amount 
of sulphur present in a furnace charge, the more FeS will there be in 
the matte after melting, and the smaller will be the proportion of 
copper. In consequence, the grade of the matte will be lower. 

The proportion of sulphur in the charge thus controls the con- 
<^entration of the copper by the smelting operation, and, in order to 
effect the desired concentration, oxygen is required in order to burn 
off sulphur and to oxidise iron. There are two general methods 
of supplying this necessary oxygen. 

(1) By a preliminary oxidation of the charge outside the smelting 
furnace — Boasting. 

(2) By oxidation inside the smelting furnace itself — The pyritic 
principle (to be considered later). 

Modern Practice as regards Roasting, — In modern copper 
smelting, the tendency is to do away with roasting as much as possible. 

Objections to Roasting. — (1) Expense involved by a separate 
preliminary process. This includes 

(a) Preparation of the ore for roasting. 

(h) Extra ground, and plant required for handling. 

(c) Labour, fuel, etc., required. 

{d) Extra handling of material before and after roasting. 

(2) Heavy mechanical and other losses during the process. 

(3) Loss of the fuel value of the iron and sulphur for smelting, 
(i) Necessity, in the majority of cases, of having the ore in a fine 

state of division in order to conduct efficient roasting, thus militating 
against its subsequent use in the blast furnace, unless the product 
receives preliminary agglomeration. 

Thus at Tennessee, the cost of roasting was about 40 cents, or 
Is. 8d. per ton of ore (equivalent to Jd. on every pound of copper 
produced). The cost for the year 1903 amounted to £19,000, 
employing 170 men out of a total staff of 900 at mines and smelters. 
The conditions for roasting were here exceptionally favourable. The 
closing of the roast-yards set at libert}^ £34,000. which had been tied 
up in this manner. 

Advantages of i^oas^m^f.— Illustrative of the conditions under 

* Though these actions represent with fair accuracy what occurs on smelting, there will 
be indicated later, in the proper place, some modification, due to interactions of certain 
oxides and sulphides, in the furnace. * 



^^llich roasting is advantageously conducted in modern practice, the 
case of the Butte second-class ores may be quoted. 

These ores contain about 5 per cent, of copper in the form of sul- 
l^liides. finely disseminated through large quantities of siliceous gangue. 
Direct smelting in a blast furnace would not yield a matte of the desired 
■ converter '' grade, except at very heavy expense and difficulty. The 
ore is, therefore, wet-dressed up to 9 to 10 per cent, copper, and the 
coarse concentrates now help to yield a good matte, when smelted in 
the blast furnace. By the wet-dressing treatment, however, a con- 
siderable quantity of fine material is unavoidably produced, for which 
the most convenient treatment in such large quantities, under prevail- 
ing conditions, is in the reverberatory furnace. The atmosphere of 
this tvpe of furnace being to a great extent neutral, the charge would 
tend simply to melt down without very much reduction of sulphur, 
resulting in the production of very low-grade matte. Roasting of these 
fine concentrates is, therefore, desirable for reducing the sulphur to 
such an extent as will yield a high-grade converter matte.* Roasting 
being thus often advisable as a preliminary, its inclusion in a smelting 
scheme under suitable conditions entails the following advantages 
over the direct reverberatory treatment of unroasted ores : — 

(1) It ensures satisfactory concentration on smelting. 

(2) It leaves reverberatory furnace smelting practically a re- 
melting operation, and so affords exact control of the concentration 

(3) The roaster gases may be utilised for making acid. 

In modern practice the work of the reverberatory plant is con- 
trolled at the roasters. The reverberatory foreman smelts whatever 
mixture is sent from the roasting plant, and if the grade of the resulting 
matte is not satisfactory, it is in the roasting operations that the 
required change is made for the correct adjustment of the sulphur and 
for controlling the consequent tenor of the matte. 

The Reactions of Roasting. — The operation of roasting is the 
\ posing of a substance to the effects of heat and air, in order to 
ddise it, and to render it more suitable for subsequent smelting 
aerations. I 

In the case of the ordinary sulphide copper ores, roasting not only 

reduces sulphur, and so ensures good concentration on smelting, 

hnt (h) by oxidising the iron, pro\ide8 a ready flux for siliceous gangues. 

* It might also be possible to assist the concentration in the inatto by the addition of 
ly available oxidised ores or residues. 

tThe chloridising roasting of copper ores is also sometimes employed in connection 
wet processes. 



The more important reactions occurring to the usual constituents of 
the copper ores which are roasted, may be summarised as follows : — 

Iron Pyrites. — First loses free sulphur at a low temperature : it 
is generally assumed that FeS is left, but the residual sulphide rarely 
attains this composition — 

FeS^ -> FeS + S. 

Iron Sulphide. — Sulphur has a great affinity for oxygen, to form 
SO9, and it may be assumed that this reaction first takes place thus — 

FeS + 0, -> (Fe) + SO, (i.) 

The iron is however instantly oxidised by the excess oxygen always 
present — 

(Fe) +0->FeO(?) (ii.) 

Or, combining (i.) and (ii.) — 

FeS + 3, -> FeO + SO2. 

This sulphur oxidation is an important source of heat, and in the 
early stages of roasting, sulphur is seen burning with the familiar 
blue flame, and the mass becomes red hot ; stirring being required to 
prevent the material from sintering by the heat generated within 

The oxidation of the iron generally proceeds further, yielding 
higher and more stable oxides — 

J 2FeO + O -> FePg. 
( 3FeO + -> FegO^. 

The SO, in the presence of oxygen and in contact with strong^ heated 
material further tends to form SO3, which is a powerful oxidising 
agent, and plays a considerable part in the various oxidising reactions 
which occur. 

Pyrrhottite behaves in much the same way ; it may be regarded 
as consisting of xFeS + a little extra sulphur. It does not roast quite 
so easily as pyrites, partly on account of physical characteristics, and 
partly because, in the case of pyrites, the greater amount of excess 
sulphur which is first driven off, tends to leave the mass more porous 
and so assists oxidation. 

Copper Sulphide. — Its characteristics on oxidation have already 
been indicated in Lecture III., p. 36, It melts easily, often at roasting 
temperatures, hence careful heating and attention are required when 
much is present. 




The reactions are probably analogous to those of FeS oxidation, 
in the primary oxidation of the sulphur and the instantaneous oxida- 
tion of the nascent copper— 

f Cu^S + 0., -> (2Cu) + SO, 
t (2Cu) + d -> Cu^O, 

thus Cu,S + 3 . -> Cu^O + SO., ; 

this being accompanied by simultaneous action of the following 
nature : — 

Cu,0 + SO, + 2, -> 2CuO + SO3 

Cub + SOg"'-^ CuSO^ 

CuSO, + Cu,0 -> 3CuO + SO.,. 

In addition to the tendency to melt, copper sulphide roasts less per- 
fectly than the FeS, usually yielding oxides which are accompanied 
by small quantities of sulphate. 

Chalcopyrite is the commonest copper ore, and the material most 
frequently subjected to roasting in copper smelting practice. 

Consisting of Cu^S . Fe^Sg, and accompanied usually b}^ a large 
excess of FeS.„ it behaves very much like a mixture of these sulphides 
when treated in the roaster furnace, hence the reactions on roasting 

II follow on the lines just indicated. 
In practice the roasting is never carried to such a degree that all the 
I sulphur is eliminated, since it is essential to retain some sulphur in 
I order to collect the copper in the form of matte, and also because 
I the time, and the cost of the fuel required to roast all of it off, would 
1 be prohibitive. Consequently, the products from the roasting of 
'^alcopyrite consist principallv of oxides of iron and copper, together 
ii a certain amount of copper sulphate, very little iron sulphate, 
■■.m\ some undecomposed sulphides. 

The actual form in which the sulphur is present at the end of the 
isting operation is not usually of very special importance in practice, 
lecially where the previous experience with the roasted material 
•ermines the extent to which the roasting is conducted, since the 
ater part of the sulphur eventually produces the sulphide and 
istitutes the matte, on smelting the roasted charge; although some 
also eliminated as SO, by interaction with oxides. In modern 
i.-iting practice, therefore, all that is usually required is to roast 
• ore down to, say, 5 per cent., 6 per cent., 8 per cent., or whatever 
'ioportion of sulphur is necessary to yield the required grade of 

1>onverter-matte in the reverberatories, as judged by previous experi- 
mce of the furnace plant and working. Much SO, is evolved during 



the roasting, though it is usually largely diluted with nitrogen from 
the air used up. 

Other Foreign Constituents of Copper Ores — Zinc Sulphide. — ZnS is 
sometimes present. Some remains unchanged on roasting, as the heat 
in ordinary practice is not great enough to thoroughly decompose it. 
Some oxide and some sulphate are also produced. 

2ZnS + 7, O -> ZnO + ZnSO^ + SO.^ 

is suggested by Peters as a probable reaction occurring to this material 
under roasting conditions. 

Lead Sidphide is also occasionally present with copper ores. It 
melts readily, and is not entirely decomposed at the temperatures 
employed for the roasting of copper ores. The reactions on oxidatioi 
are largely analogous to those for other sulphides. 

PbS + O. -> Pb + SO. 
Pb + -> PbO 
or, PbS + 3 . O -> PbO + SO2. 

Also, 2PbO + SO3 -> PbSO^ . PbO (basic sulphate)} 


Arsenides are partly left as the corresponding oxides, whilst som< 

As^Og is evolved, and some basic arsenate generally remains. 

Roasting Practice. 
Favourable Conditimis for Successful Boasting. 

(a) The sulphide should be in a finely divided form, so as to ensure 
good contact with the air. 

(b) The air should be supplied in a gentle current, so as to con- 
tinually provide fresh oxygen, and sweep away the inert gases which 
are produced. 

(c) The ore should be heated to a dull red heat, which is a con-; 
dition favourable for commencing the ignition and reactions. The 
temperature should, of course, be well below a melting heat (Peters). 

The Apparatus for Roasting depends to some extent on the clase 
of material to be dealt with, which may be in the form of either (a 
lump ores, or (b) fine ores. 

(a) Roasting of Lump Ores. — Tn modern copper-smelting work, th< 
practice of roasting lump ores is practically obsolete. The con 
ditions under which its use might still be justified are those associ 
ated with newer mining districts, where rapid concentration of heav; 
sulphide ore into matte is required, before the time is ripe fc 



smelting the material pyritically, and where further, it is desired to 
employ the blast furnace for the smelting operations under these 

The advantages possessed by the method are — 

(1) No preliminary crushing is required. 

(2) The product is largely in the form of lumps, and hence imme- 
diately suitable for blast-furnace work. 

(8) The plant and appliances required are simple. 
The two methods employed are — (A) open-air roasting, (B) roasting 
in kilns. 

A. Open-air Roasting of Lump Ores. — This method is conducted 
in heaps or stalls, and the features just considered apply particularly 
to this branch of roasting practice. The modern tendency is to avoid 
heap-roasting altogether, and it is only conducted when the con- 
ditions are exceptional. 

Amongst the many grave objections to open-air roasting are — 
ia) It is very slow, since a long period of time is required for the 
oxidising effect to penetrate through massive lumps of ore. 

(h) A large amount of capital is tied up in the material at the 

(c) The losses occasioned by wind and rain are very considerable. 
{d) It is difficult to use up a large quantity of fines in the roast- 

(e) Difficulties arise owing to damage by the fume, and from inter- 
lence b}'^ litigation. 
There is one special instance of a modern smelter making a great 
cess of heap-roasting — namely, at Rio Tinto — but the circum- 
nces are peculiar, as the roasting is followed by leaching operations 
the immense ore heaps in situ. 

This branch of roasting need not be considered at length, and the 
ler standard text-books give full descriptions of the various 
' thods employed. The following particulars are important, how- 
f-r, when under exceptional circumstances such work has to be 
iflertaken : — 

The maximum and best average size under ordinary conditions is 

' feet by 24 feet, by 7 feet high above the bed of fuel. The height is 

portant, and varies with the quantity of sulphur in the ore. The 

ver the sulphur content, the higher the pile ; with about 40 per cent. 

iphur, the best height is 6 to 7 feet ; with 15 per cent, of sulphur, 

to 9 feet ; and if still less sulphur be present, the height may 

be a little greater. Such a heap holds about 240 tons, and if 

quantity of ore to be dealt with exceeds this, a number of such 


piles should be constructed. The time occupied in roasting is about 
70 days, with 10 days more for removing and rebuilding. 
The selection of a proper site is important. 

(a) The prevailing direction of the wind must be considered, so 
as to keep the fumes away from the works and offices. 

(h) The yards must be protected from winds, so as to prevent 
losses of dust, as well as uneven burning. 

(c) The ground must be perfectly dry or drained. 

Along the upper edges of the roast-yard a deep trench should be 
cut, so as to catch rain-water, and prevent it from washing soluble copper 
salts out of the pile ; drainage trenches must also be provided to carry 
any copper-bearing liquors to some point where the copper can con- 
veniently be precipitated on scrap iron. Enormous losses of coppei 
may occur if these precautions are not observed ; thus, at one period 
the old roasting process in Tennessee, as much as 34 per cent, of thej 
copper in the heaps was lost in 186 days. 

Preparing of the Floor. — Remove roots and subsoil, fill space witl 
broken stone or rough tailings, cover with 4 to 6 inches of clayey loam, 
and beat down well. The floor is then fairly impervious, and doeg 
not crack on drying. The ground should be given a gentle slope so as' 
to facilitate draining. A layer (about 6 inches thick) of fine ore is next 
put down, then 9 inches of fuel ; channels are now mapped out by 
means of logs set in both directions, leading to rough chimneys. The 
pile is then constructed, with the lower parts of the very coarse 
materials, smaller stuff being put towards the top and sides. On the 
very top and at the outside of the pile are placed the fines, but this 
top cover is only put on when the burning is well started. This 
process is still worked at Tyee, B.C., and at some other localities, but 
is most probably only a temporary plan, to be replaced by a more 
efficient method as development progresses. 

B. Kilns for Lump Ores. — Kilns possess the advantage that they 
permit of arrangements being made for the recovery of SO2 for acid 
manufacture, and the subject belongs more properly to that branch 
of technology. Few large smelting works employ kilns for roasting 
lump ores, though there are important exceptions at works both in 
Britain and on the Continent of Europe. Kilns are used at the Cape 
Copper Company's smelter at Britton Ferry, for this purpose. 

(b) The Roasting of Fines.— Fines (and particularly fine concen- 
trates) are the usual materials subjected to roasting. The finer the 
particles, the more rapid and complete is the oxidation, but the losses 
by dust are heavier. The size limit is thus liable to some variation 
but often the material roasted is that under 1-inch in size. 


Roasting Furnaces — Requirements. — For the roasting of fines there 
is simply required a place where the material can be gently heated in 
the presence of a constantly renewed air supply. The fuel has itself 
a reducing action, it must therefore be separated from the charge, 
and hence the furnace employed is of the reverberatory type. Muffles 
are never used for the oxidising roasting of copper ores. Since only 
a moderate temperature is necessary for the operation, the furnace 
needs but a small fireplace, and it is provided with a large hearth 
area. The fuel used is one yielding the fairly long oxidising flame 

Developments of Roasting Practice. — The main objects sought in 
roasting practice have been — 

(1) To have as large a surface of material exposed to heat and air 
as possible. 

(a) By elongating and multiplying the beds of the furnace. 
(h) By furrowing and rabbling the charge. 

(2) Continually to expose fresh surfaces of ore to oxidation. 
(a) First by hand-rabbhng. 

(6) Later by movable furnace hearths, 
(c) By mechanical rabbling. 

(3) To obtain a continuous output — 

{a) By mechanical charging, rabbling, and discharging. 

The Development of the Roasting Furnace. 

A. Fixed Hearth. — In Great Britain from 1583 onward, roasting 

in small reverberatory furnaces seems to have been the usual method, 

and up to 1850 the furnaces appear to have been only of moderate 

dimensions, with a single hearth, 16 feet x 13 feet 6 inches, constructed 

I of firebricks set on end, and with a fire-box 7 feet x 2 feet 3 inches 

X 18 inches. Rabbling was done by a long rake, the material being 

charged and worked through one door. This method of working 

wasted time, made the process intermittent, and caused continual 

cooling down of the furnace, involving large fuel costs and much 

hour. The first improvements were to lengthen the hearth, to add 

more working doors, and to put the charge into the furnace by a 

hopper passing through the roof. It was next found best to elongate 

the hearth still further, and to drop the level of the bed in stages by 

Nout 2 inches at a time, thus ensuring better control of working. 

iv this means the best type of hand-calciner was arrived at, con- 

fing of four beds, each 16 feet x 16 feet, the whole charge being 



Id roasting, the ore is first placed in the coolest part of the 
furnace, and is worked towards the fire, so that the charge travels 
in one direction, and the flame and furnace gases in the opposite 
direction to meet it. 

The advantages of this system are that — 

{a) The clotting of the sulphides is prevented, since the first part 
of the roasting proceeds at a comparatively very low temperature. 

(h) The sulphur in the ore often provides sufficient heat to 
maintain the roasting in progress during the early stages. 

(c) The hottest parts of the furnace are where the roasted infusible 
oxides arrive, so but little clotting or sintering occurs here. 

The capacity of the four-bedded hand-roaster is 7 to 15 tons 
per twenty-four hours, depending on the sulphur proportion in the 
charge and in the roasted product. 

It is a very useful form of furnace when labour is cheap. The 
furnace works very efficiently, but in the New World, where manual 
labour was dear, labour costs became prohibitive, and in order to 
economise in this direction, mechanical rabbling was introduced. 

The O'Harra Calciner (1885) was essentially the old type of furnace, 
double hearthed and mechanically rabbled. It consisted of long 
straight furnace hearths. The rabbles were ploughs dragged through 
the furnace by means of endless chains which were carried over 
grooved pulleys, situated outside the furnace, at the ends. This 
was an important invention, giving a continuous feed and dis- 
charge, a much larger output, and efficient and regular stirring 
without much hand labour. The rabbles became cooled on issuing 
from the hearth. The capacity was 50 tons per day from furnaces 
of 90 feet x 9 feet hearths, giving a roasting capacity of 61 lbs. 
of ore per square foot of hearth area, compared with about 33 lbs. 
per square foot with the old hand calciner. In working the early 
forms of this furnace there were many mechanical troubles and break- 
downs, and the subsequent modifications of this form consisted largely 
of devices for the purpose of overcoming such difficulties 

Modificatio7is and Improvements. — Allen, instead of a rope to carry 
the ploughs, used small wheeled carriages, running on a track which 
was laid along the floor. 

Broivn (important) ran the carriages along narrow corridors at 
either side of the hearth, so as to protect the ropes and carriages from 
the very corrosive action of the furnace gases. A continuous narrow 
slit along the inner wall of the corridors allowed the arm carrying the 
plough to travel forward. 




Wethey : Keller ; worked on very similar principles. The chief im- 
provements were in details, and had for their object the prevention 
of wear and tear, and of the break-down of parts. 

Prosser. — Very similar ; used at Swansea Works. 

Ropp. — The carriage rims underneath the bed, and supports a 
vertical shaft which passes through a slot along the furnace hearth 
and carries the arms furnished with ploughs. 

Fig. 17. — O'Harra Furnace (Fraser-Chalmers), illustrating Principle of Mechanical 
Rabbling by Travelling Ploughs. 

The Ropp and Prosser calciners work very successfully. The hearth 
\> about 105 feet long x 11 feet wide, with a capacity of about 36 tons 
per day. 

Fig. 18. — Section through Mechanically Rabbled Roaster Furnace (illustrating 
Improvements for Protecting Driving Mechanism). 

Brown Horse-Shoe Furnace operates on the same principle as the 

>ove, except that the hearth is bent round in order to save space. 

Pearse-Tnrrett (1892 at Argo). — In this type of furnace the bed 

curved round in the form of a circle. The rabbling ploughs are 

irried at the ends of arms which are attached to an upright rotating 



spindle. The spindle is set in the centre of the space enclosed by 
the circular hearth. 

In all the above classes of furnace, the firing is done, when 
necessary, from fireplaces built at intervals along the sides of the 
furnace ; either coal or gas being emploj^ed as fuel . 

B. Rotating Hearths. — This type of furnace is still reverbera- 
tory, but instead of making use of mechanical rabbling, the hearth 
rotates, in order to give agitation to the materials and assist their 

(a) Intermittent Working — The Briiclaier Roaster. — The details and 
working of this roaster are familiar. The furnace was invented in 
1864 for gold and silver ore-roasting in Colorado, and was later intro- 
duced for the roasting of copper ores, being at one time the furnace 
most commonly used for the purpose. It was employed all over the 
Western States, and at one works alone, 56 were at one time in use. 

The usual length was 18 feet 6 inches and the diameter, 8 feet 
6 inches; giving an output of about 12 tons per twenty-four hours. 
It was furnished with a removable fireplace, used to start the roasting. 
The operation could then be allowed to proceed by itself, the fireplace 
being wheeled away to another hearth, and being eventually brought 
back to the first hearth for about three hours, in order to give the 
required higher finishing temperature. Several dust chambers were 
attached to this, as to all forms of roasting furnaces, which by their 
nature and manner of work are apt to produce considerable quantities 
of dust. 

The advantages of the Briickner cylinder lay largely in the fact 
that it afforded good control of the sulphur contents in the charge, 
since the ore could be retained in the furnace until the sulphur was 
sufficiently low. The furnace is simple to work, and not so liable to 
get out of order as many other forms. It possesses however, distinct 
disadvantages in that its working is intermittent, its use involves 
comparatively high fuel costs, whilst the discharging presents con- 
siderable difficulty and trouble to the labour employed, on account 
of the awkwardness and the high temperature of the discharge, and 
the sulphurous gases evolved. 

Its use has now been very largely discontinued. 

Improvements — {h) Continuous Working. — The continuous type of 
roasting furnace of this class involves the use of sloping cylindrical 
hearths which rotate, and so agitate and help to discharge the 

Oxland' (1868) first introduced this type in Cornwall, for the roasting 
of tin ores. 



The Oxlaiid furnace was an inclined cylinder, the material was 
fed in at the top, and by the rotation of the cylinder the charge gradu- 
ally travelled downwards, approaching nearer and nearer to the fire, 
and being discharged close to the fire-box. 

Whitp, (1872) improved this furnace, and the White cylinder is 
largely used in South Wales. The cylinder revolves slowly by friction 
gearing; inside are four lines of projecting brick-work which form a 
shelf, thus assisting the agitation of the charge. 

The White-Howell Furnace is somewhat similar to the White, but 
is unlined for the greater part of its length, except at the lower end 
near the fire-box, where it is much wider and is bricked. It is stated 
to Avork more satisfactorily than the older form, having a larger capacity 
and using but little fuel. 

The furnace is employed at the Cape Copper Works, South Wales, 
for matte-roasting. It is here 60 feet long, 7 feet diameter, incUned 
6 inches in 60 feet, makes 8 revolutions per hour, and has a capacity 
of 10 tons of charge per day. 

Arqall Furnace. — Consists essentially of four narrow tubes bound 
together, each 28 feet long, 2 feet diameter, and lined. It works 
rapidly, having a capacity of 40 to 50 tons per day, but is used more 
for the roasting of cupriferous gold ores than at the copper smelters. 

C. The MacDougal Type. — The most important form of modern 
)aster furnace, and that most generally employed, is the MacDougal 
rvpe. The first furnace on this principle was invented by Parkes in 
1860. The design embodied two hearths, one above the other. Verti- 
caUy down the centre of these passed a spindle, supporting arms from 
which were suspended the ploughs, and the rotation of this spindle 
carried the arms over the beds. 

As devised by Parkes, various mechanical difficulties were found, 
and the working was intermittent, but the principle was recognised as 
important. MacDougal in 1873 introduced his modification of the 
•furnace, primarily for the roasting of pyrites, at a Liverpool works, 
and this form has now supplanted many of the older types for 
copper ore roasting, and is in operation at most of the new smelting 

Princijdes of the MacDougal Type. — The furnace consists of an iron 
ylinder lined with brick. Six circular hearths are constructed inside, 
e above the other, and the vertical spindle carrying the arms and 

ughs for each hearth passes through the centre of the furnace. 

e ore is ploughed towards openings on each hearth, which com- 

unicate with the hearth below ; the charge thus travelling from the 

ter edge towards the centre, through the central opening to the 






middle of the next floor, then outwards to the openings at the 
edge, and so on. The original MacDougal furnace was J 2 feet high 
and 6 feet in diameter. It was improved by Herreshof in the 
direction of better rabbling mechanism and greater ease of repair. 
The central spindle was an air-cooled shaft; the supporting arms 
were made so as to be easih' removable from the shaft to facilitate 
repairs, and the furnace was enlarged. Herreshof used air-cooling 
for the spindle and arms, as shown in Fig. 20. 

N9I Hearth 
N92 Hearth 
N94 Hearth 

Fig. 19.— MacDougal Ivoaster — 
Vertical Section. 

Fig. 20. — Herreshof Furnace — Section indicating 
Connections for cooling Rabbles and Spindles. 

Evans, and subsequently Klepetko, in working the furnaces in 
Montana, introduced, in about 1892, various marked improvements. 
The dimensions were increased, enlarging the output. The spindle 
and arms were water-cooled, which improvement removed much 
of the great difficulty in working the MacDougal furnace, Avhere the 
rapid wearing out of working parts, and the difficulty of their removal, 
repair, and renewal interfered greatly with efficient A^orking. 



]\Iany of these troubles have now been overcome in the Evans - 
Klepetko t^'pe; and in the still farther improvements since made at 
Anaconda. The general arrangement of the floors, spindles, arms 
and other details shown in the Herreshof furnace (Fig. 20) are 
preserved in the Evans-Klepetko and similar types of roaster ; the 
chief alterations are in matters of detail, the results of which have 
however, been important. 

Furnaces of this improved kind are now used all over the West ; 
there are 64 at Anaconda, Mont. ; 32 at the International Smelter, 
Tooele, Utah ; 24 at Garfield, Utah ; 16 at Steptoe, Nevada ; and 
also at Balakala, CaL, Cerro de Pasco, Peru, and other large smelting 

Important Advantages. — Of the marked advantages of this type of 
furnace, the following are perhaps the most striking and important : — 

(1) There is a great saving of floor space by having the six 
hearths one above the other. 

(2) The use of a central common spindle carrying the arms and 
jjloughs simplifies the mechanism. 

(3) The form is convenient for the compact arrangement of a 
oasting plant of many units for feeding, discharge, and supervision. 

(4) Very little heat is lost by radiation, as the heat passes mostly 
rom one hearth to another. 

(5) Very Httle fuel is required, none with heavy sulphides (except 
ior starting), as the heat of oxidation of the iron and sulphur usually 
yields a high enough temperature to keep the operation going. The 
fuel costs are lower than in other types of roaster. 

(6) Thorough rabbling, greater uniformity and better mixing of 
product, continuous and regular feed and discharge are obtained. 

(7) The roasting is thorough, and perfect control of the degree of 
oxidation is ensured b}^ adjusting the rate of passage of the ore through 
the furnace, which is regulated by varying the ore feed and the speed 

f rotation of the rabbles. 

(8^ Great saving in labour costs and difficulties. The labour in 
oasting plants is extremely arduous, on account of the high 
' mperature of the material, and is dangerous on account of the 

The Evans-KIepetko-MacDougal Roasting Furnace Plant 
it Anaconda* — The roasting plant at Anaconda formerly consisted of 
">r» Briickner cylinders, which were eventually all scrapped and replaced 
by new plant of the MacDougal type, subsequently greatly modified 
and improved as one difficulty after another had to be overcome. 

The saving in working costs resulting from this replacement of the 



Briickners by MacDougal roasters is reckoned at about 5 cents (2|d.) 
on every ton of calcines treated. 

The roasters are arranged in four rows of 16 each, running east 
and west. The charge cars travel along tracks at a height of 20 feet 
above, discharging into rows of bins, one situated over each calciner. 

Deiails oj Furnace. — Height, 18 feet 3 J inches ; diameter. 16 feet. 
Six hearths. The spindle is made in three lengths, each to carry the 
arms for two hearths ; it is 18 inches in diameter and is water-cooled. 
The rabble arms are 6 feet long, half round, and flanged on the lower 
side ; they too, are hollow and water-cooled. The rabble-blades 
were formerly cast in one piece with base plate, so as to slide on to 
the arms, but are made now with detachable blades, which slide into 
grooves on the base plate, so as to facilitate removal for repairs ; the 
blades are 6 inches square and IJ inches thick (Fig. 32). 

The arms on separate floors are set 
alternately at right angles. Of the two 
arms for each floor, one carries six blades, 
the other seven, so that the furrows result- 

Bl IIIKl O^^^^^^s^ ^^^ from one set of blades are turned over 
^1! IIIM il ^ =^^ \yy the other. The blades are set so as to 

direct the ore from the outer to the inner 
edge or vice versa, according to the parti- 

HJIHl I llf^ cular hearth. The spindle and connections 

lllfll j 11 ^ are protected from falling ore by shields 

which are bolted on. The rabbles move 
slowly, making a 2|-inch furrow in a 5-inch 
layer of material. 

Capacity. — 40 to 45 tons per day each, 
reducing the sulphur in the charge from 
^'S- 21. -Spindle Connections 30 per Cent, to about 8-0 per cent. The 

and Guide Shields of Evans- . . r j.i i j. • t_ j. o r^rxrx j. 

Klepetko Roasters. Output of the plant IS about 3,000 tons 

of calcines daily. 
System of Working. — Since reverberatory furnaces are used essen- 
tially as remelting furnaces only, the roasting plant is operated so as 
to yield a product of such composition as will directly produce a suitable 
matte and slag on melting in the reverberatories. The fluxes required 
for the calculated reverberatory charge are, therefore, sent through the 
roasters mixed with the fine concentrate ; such practice possessing 
many advantages. The charge thus consists of fine concentrate from 
the concentrator setthng tanks, and screened lime-rock flux (too fine 
to be used in the blast furnaces). The limestone hghtens the charge, 
decreases the tendency to clotting of the pure sulphides, chemically 



assists oxidation, preheats and thoroughly mixes the flux, and ensures 
a uniformly mixed charge for the reverberatory furnaces ; whilst the 
extra cost involved is but very small. 

Three per cent, of lime is used ; 40 tons of concentrates, 1 J: tons of 
lime-rock, and IJ tons of flue -dust being charged per twenty-four 
hours per furnace, through an automatic gravity feed, the opening of 
which is closed and opened by an eccentric. The speed of the eccentric 
and the extent of the opening are adjustable. 

Working. — Charge contains 25 to 35 per cent, sulphur. 

1st Hearth.— Temper a,ture about 230° C. (black heat). This is 




' — 



Fig. 22. — Kabble-blades and Bases. 

practically a drying floor, and the wet ore wears the rabbles away 
rather quickly. Special forms of plough are being introduced. About 
4 per cent, of sulphur is driven off from the pyrites. 

2nd Hearth. — Hotter ; not quite red, except near outer edge. 
About 5 per cent, of sulphur burnt off. 

3/yZ Hearth.— Bright red heat (about 700° C). Sulphur can be 
seen burning off the ridges of calcines, with a blue flame. 5 per cent, 
of sulphur eliminated. There is some clotting, and the sinter sticks 
to the rabble-blades, and has to be barred off occasionally. 



Uh Hearth. — Bright red heat (about 750° C), uniformly bright, 
but the flame has ceased. Sulphur loss, 4 per cent. 

mil Hearth.— The hottest (800° C). Bright red. 

Bottom Hearth. — Cooler, dark red (about 650° C). The doors on 
this floor are left open. The charge is guided towards openings at the 
outer edge to discharge chutes whilst still red hot, and it is fed from 
here whilst hot into the reverberatory furnace-bins. 

Efficient dust catchers and settlers are essential on the roasting 
plant. The gases escaping at a temperature of about 315'^ C. contain 
2 per cent, of SO^ by volume, 5 per cent, by weight. The ore takes 
2| hours to pass through the furnace. Practically no fuel is required 
except to warm up the roaster on commencing work. 

Labour. — The requirements are small. There is one general fore- 
man for the plant, and two helpers for each set of four furnaces. The 
conditions are rather trying, especially during the discharge of the 
calcines into the reverberatory charge cars. 

Roasting Ores poorer in Sulphur, in MacDougal Roasters. — 
The Anaconda concentrates carry sufficient sulphur (33 per cent.) to 
supply all the heat necessary for carrying out the roasting operations. 
When the sulphur is below this requisite quantity, some extra heating 
may be required, though, on the other hand, the reduction Avhich is 
necessary in the sulphur contents is lessened, depending, of course, 
on the proportions of copper and iron in the charge. At Garfield, 
Utah, where the concentrate only contains 20 per cent, of sulphur, 
the fuel required for all roaster purposes is equivalent to 0-2 per cent, 
of the charge, one of the calcines' outlets being converted into a fire- 
place. Here the output per furnace per day approaches 55 tons, 
roasting the sulphur from 20 per cent, down to 10 to II per cent. 
The flue-dust losses at this plant are 6 per cent., so efficient dust 
catching appliances are essential. 

The Costs of Roasting in the MacDougal Furnace, — Ricketts 
has recently published a valuable analysis of the costs of the 
roasting operations at the Cananea Smelter. The figures must, 
however, be understood to apply strictly to the conditions prevaiHng 
at this particular camp. 

The roaster plant consists of 32 improved MacDougal furnaces. 
The charge supplied to the roasters assays — 

Copper, . . . . .5-2 per cent. 

Iron, . 
Silica, . 







j whilst the product (*' calcmes '') has an average composition of 


6-3 per cent 

Iron, . 

. 345 „ 


- 7-7 „ 

Silica, . 

. 28-6 „ 


. 44 „ 

The plant operated on the following quantities of material, from 
February to July, 1911, inclusive : — 

Concentrates, . 
Fine sulphide ores, . 

Total charge, . 
Weight of " calcines " pro- 


32,929 short tons = 76-08 per cent, of charge. 
9,590 „ - 22-16 



= 1-76 

- 10000 
= 82-10 

- 17-90 

The total costs of roasting (from roaster charge-bins to reverberatory 
furnace) worked out at 38-45 cents per ton, the distribution of these 
costs being as follows : — 

Bedding, . 
Operating furnaces, . 
Hauling calcines. 
General expenses, 

Total Costs. 

Cost per Dry Ton. 


Total direct costs. 
Cost of flux. 

. §15,543-38 


Total costs, 

\nalysis of Cost — 
(1) Operating — 
Labour, . 
Power, . 

Water, . 

. §16,641-23 

. §7,398-56 








(2) Repairs— 

Labour, . 
Shop expense, . 
Supplies, . 


. $2,507-35 





Total costs, . 

. §16,641-23 




Peters, E. D., "Principles" and "Practice of Modern Copper Smelting." 

Cloud, T. C, "The M'Murty-Rogers Process for Desulphurising Copper Ores." Trans. 

Inst. Min. and Met., vol. xvi., 1906-7, p. 311. 
Hofman, H. 0., "Recent Progress in Blast Roasting." Bulletin Amtr. Inst. Min. Enr,., 

No. 42, June, 1910. 
Austin, L. S., "The Washoe Plant of the Anaconda Copper Mining Company." Trans. 

Amtr. Inst. Min. Eng., vol. xxxviii., 1906, p. 560. 
Rickets, L. D., "Developments in Cananea Practice." Engineering and Mining Journal, 

Oct. 7th, 1911, p. 693. 
Redick F. Moore, " Recent Reverberatory Smelting Practice." Engineering and Mining 

Journal, May 14th, 1910, p. 1021. i 

See also — 

Pulsifer, H. B., "Important Factors in Blast Roasting." Met. and Chem. Eng., 
1912, vol. X., No. 3, March, pp. 153-159. (With good Bibliography.) 

Editorial Correspondence, "Sinter-Roasting with Dwight-Lloyd Machines at Salida, 
Col." Ibid., 1912, vol. X., No. 2, Feb., p. 87. 

Dwight, A. S., "Efficiency in Ore-Roasting." School of Mines Quarterly, 1911, 
vol. xxxiii.. No. 1, Nov., pp. 1-17. 

8i V 



Reverberatory Swieltinc Practice. 

Functions of the Reverberatory Furnace— Requirements 
for Successful Working"— Principles of Modern Re- 
verberatory Practice— Operation of lYIodern Large 
Furnaces— Fuels for Reverberatory Work; Oil Fuel; 
Analysis of Costs— Condition of the Charge. 

The Functions of the Reverberatory Furnace, — The reverber- 
atory is essentially the furnace for the smelting of fine material, as 
the comparatively still atmosphere, the absence of blast, and the 
opportunities for settling prevent the heavy losses by dust which 
necessarily accrue with the other types of smelting furnace. The 
atmosphere of the furnace is practically neutral, it therefore exercises 
little influence on the reactions taking place in the charge, and the 
reverberatory is, in consequence, mainly a melting furnace. 
Its functions are : — 

(a) To allow of the formation, from the mixture of sulphides and 
oxides in the roasted materials from the calciners, of a copper matte 
and a slag. 

{b) To maintain such a high temperature as to render these products 
perfectly fluid, and thus to allow the matte and slag to settle and 
f'parate thoroughly. 

In spite of the neutral atmosphere, however, the smelting of the 

roasted materials usually results in a higher concentration than wduld 

be expected from the calculation of the sulphur, copper, and iron in 

the charge. The reason of this is that the smelting operation results 

1 some further elimination of the sulphur, which causes the produc- 

on of a higher grade matte. This additional elimination of sulphur 

J the reverberatory furnace smelting of the roasted charge is due 

To the reactions which take place on melting, between the oxides, 

sulphates, and sulphides of copper, all of which exist in the products 

from the roasters. These reactions are expressed by the equations — 

CU2S + 2CU2O -> 6Cu + SO, 
CugS + CuSO^ -> 3Cu -f 2SO2, 
which indicate a further addition of copper to the matte, and a 




Thus a typical reverberatory charge 

corresponding loss of sulphur, 
of the following composition :- 

Iron, . 

should theoretically yield, on melting down, a matte running- 

*Cu(8-3) ->Cu,S 10-4^ r Cu 8-3 

= S 8-4 
S (8-4-21) -> FeS 17-6 j I Fe 11-3 

In actual practice however, the matte resulting from the reverbera- 
tory smelting of the charge had the composition- 

j Cu 45 per cent. 

' S 27 

27-2 per cent. 

^ a ,, 

8-4 „ 

f Cu 30 per cent, 
or J S 




Fe 28 

the 3 per cent, loss of sulphur causing a 15 per cent, increase in the 
copper contents of the matte. 

Experience in the working of the plant enables the management 
to determine this important factor with fair accuracy, and thus from i 
knowledge of the composition of the roaster product, to regulate anc 
control the grade of the matte produced at the reverbcratories. Ii 
modern reverberatory practice, therefore, the control of the furnace 
products is carried out at the roasting plant, and the reverberatory 
furnace has simply to melt the charge and ensure good settling. 

Anaconda Practice affords a good illustration. The foreman of 
the reverberatory furnaces simpty charges what is sent him from the 
roasters, and practically nothing else is put in, I his duty being to smelt 
this mixture and to obtain from it a clean slag and fluid matte. He 
is not responsible for the grade of the matte, and if this is not satis- 
factory, some change is made in the working at the roasters. The 
reverberatory foreman does not learn the composition of the materials! 
passing into his furnace until he is furnished with the daily assay 
reports on the following da3^ 

Reverberatory smelting is essentially a British process, developed 
in Wales, as already explained, owing to a plentiful supply of good 

* When copper combines with sulphur. 
2,Cu:Cu2S::2 x 635 

■j" Some flue-dust is also melted down. 

When sulphur combines with iron. 

2 X 63-5 + 32 

S: FeS:: 32: 56 + 32 


::32: 88 

5 approximately. 

:: 4: 11. 


furnace coal yielding a long flame, and also of good refractory material. 
Many Swansea workmen were, in the early days of American develop- 
ment, and are still, employed in charge of such copper furnaces, and 
it is largelj^ due to British technical skill and to American genius for 
organisation and development that reverberatory smelting in the 
large furnaces at modern works has become so very successful. 

The Principles of Modern Practice. — Success in modern rever- 
beratory work has been due to the recognition of the fact, that with the 
maintenance of constant high temperature on large masses of material, 
thorough fusion and separation of the products can be very efficiently 

The Requirements for Successful Reverberatory "Work »— Since 
the action in the furnace is performed mainly by the effects of heat, it 
is necessary that — 

A. The melting should be as rapid as possible. 

B. The losses of heat during melting should be reduced 

to a minimum. 
The temperature required for the formation of slag and for obtaining ■ 
a thorough fluidity of the materials is from 1,400° to 1,600° C, and 
I the methods of achieving the proper conditions can best be stated as 
il the avoiding of all circumstances likely to cool the furnace or to 
' interfere with the melting down of the charge. 

A. To ensure rapidity of melting, it is essential that a very large 
<|tiantity of coal shall be burned as rapidly as possible. This requires — 

i. A large grate area, 
ii. A good draft. 

iii. The firing and grating to be conducted so as to interfere as 
little as possible with the regularity and degree of heating. 

In localities where a supply of suitable coal is not available, other 
"thods of heating, such as the use of oil or gaseous fuel, are necessary. 

B. To prevent heat losses as much as possible, it is necessary — 

i. To avpid leakages of cold air into the furnace. 

ii. To prevent radiation of heat through walls and roof, 
iii. To prevent the hearth from being cooled by the with- 
drawal of heated charges and the substitution of fresh 
and cold ones. 

iv. To utilise the heat of the already melted charge for the 
heating up of the fresh ore. 

V. To avoid as much as possible, waste of heat by the 
escaping gases. 


A. For Rapidity of Melting. 

A. i. — Enlarged Grate Area. — In the older methods of working, there 
was a general tendency to employ a furnace of standard size, anc 
improvements in the economy of the process were in the direction oi 
reducing the fuel bill as much as possible for the given size of furnace 
This was effected by keeping the grate area fairly small. 

In modern practice, economical working still involves having the' 
ratio of size of hearth to size of fire-box as large as possible, but 
instead of reducing the dimensions of the fire-grate to suit the 
hearth, a large grate is built to commence with, and the hearth is 
constructed of such a size as will utihse all the heat available. From 
this principle of burning a large quantity of fuel and melting with it 
as much charge as possible, the efficient and economical working of 
large furnaces has been developed. 

A grate area of about 28 square feet is now regarded as the minimum 
for economical work at modern smelters, and fire-boxes up to 128 
square feet in area are usual in practice. 

In small fire-boxes, only small quantities of fuel can be burned at 
once, and in consequence, fresh firing is continually required, which 
interferes greatly with the work of the furnace and decreases the 
rapidit}^ of heating. Each addition of cold fuel has a cooling effect 
on the fire and furnace gases, the temperature in the hearth being 
found to drop for a period of five or ten minutes by as much as 
100^' C, the flame becoming smoky, red, and cold. A similar time 
is required for the original temperature to be attained once more. 
Cold air is also admitted every time the fire-box doors are opened for 

The advantages of large grate area therefore include : — 

(a) Much less cooling of the furnace by frequent additions of fuel. 

(6) Higher temperatures, owing to the increased calorific intensity 
of large quantities of fuel burned at once. 

(c) Less blanketing of the fire by fuel additions. 

{d) Less chance of the whole of the grate area being clinkered up 
at once, and in consequence, less likelihood of interference with the 
rapid combustion of the fuel. 

The most rapid and economical smelting at the present day requires] 
that at least 0-7 lb. of coal be burned per minute per square foot of 
hearth area. 

A. ii. — Draft. — The charge in a reverberatory furnace hearth 
melted chiefly by the heat from the hot gases passing over it, an< 
in giving up their heat to the charge, the gases become cooled down! 



The heating of the charge is made continuous by the continual 
addition of fresh fujel in the fire-box, and by the drawing of tlie flames 
over the hearth by means of flues situated at the other end of the 
furnace and leading to the stack. The flues and stack must be large 
enough to cause sufficient draft through the furnace for the heated 
gases to be draAvn over the charge with sufficient rapidity, and much 
unsuccessful work has been due to the fact that these requirements 
have not been fulfilled. There should be a suction equivalent to at 
least 1 inch to 1-5 inches water pressure up the stack, this being 
readily measured by water-manometers — a feature of modern working. 

Reverberatories may be worked either hj forced or natural draft, 
the latter being usually preferred, though it necessitates a large stack 
and spacious flues. 

Forced draft by fan or blower under the fire-grate has been in use 
at several smelters, the ashpit then being closed. It was at one 
time adopted at Anaconda, but was given up later. The use of 
forced draft has the advantage that leakages of cold air into the 
furnace are to a large extent prevented, hot gases tending to be 
forced out rather than cold air drawn in, but the objections to its 
use include the facts that — 

ia) Special power and machinery are required. 

(b) The intense action near the fire-bars produces, from the ash 
iof the coal, a massive clinker in a semi-fused condition, difficult to deal 


(c) It is stated by smelters to have a cooling action near the fire- 

A. iii. — Firing and Grating. — This question is closely connected 
twith the dimensions of the grate, since the use of a small fire-box 
! necessitates methods of firing and grating which are not conducive 
Ij to the most rapid and efficient combustion of the fuel. In addition 
to the cooling action of frequent fresh fuel charges in the small fire- 
place, attendant disadvantages include the closing up of the spaces 
Jin the grate by which air enters for burning the fuel, and the con- 
If sequent necessity for frequent grating with small beds of fuel, which 
^ails numerous objections. 

The addition of fresh coal to the fire causes the production of large 
mtities of volatile hydrocarbons which require an increased air 
!»ply for proper combustion, and this air admission is just pre- 
•;ted by the blanketing action of the fresh fuel added. This is 
licated by the red smoky flame, and means waste and cooling. 


at the fire-box end of the furnace, near the fire-bridge, and by the 
opening of these directly after firing, the volatiles are immediately 
burnt up. This is an important feature in successful working, and 
with a large fire-grate and this air-admission, the effect of adding 
even 1-| tons of fuel on to the fire at once causes little difference in 
the furnace temperature. The flame is observed through a window 
let into the ofi-take flue, which allows of the changes in appearance 
being noted by the fireman on the fire-box platform. 

The fire is kept moderately shallow, to allow of rapid burning of 
the fuel, though deep enough to keep up the enormous body of heat 
necessary in the furnace. 

B. The Prevention of Heat Losses. 

B. i. — Avoiding Leakage of Cold Air. — The admission of cold air 
was the cause of much waste in the older processes of working. Each 
time the doors were opened, either at the fire-box. or during charging 
on to the hearth, large quantities of cold air were admitted ; air 
entered through the working door whilst slag was skimmed off, whilst 
matte was being tapped, and whilst the furnace hearth was being 
clayed ; all of which operations occupied considerable time. The 
doors were opened during the levelling down of the fresh charges, 
and at later periods when the charge was stirred and the half-fused 
masses sticking to the bottom were worked up. 

In modern practice, an essential feature of working is to keep all 
the doors closed as much as possible, and, as will be indicated shortly, 
every means is taken to eliminate the heat losses from the causes just 
referred to. Air leakage is also occasioned by bad grating, which 
causes the formation of channels in a few parts of the bed of fuel, 
admitting excess of air at these places, instead of causing it to come 
regularly through the bed in all parts. Channelling is now checked by 
the drop of suction-pressure in the flues, as registered by the manometer, 

B. ii. — Prevention of Radiation through Walls and Roof. — Such 
heat losses are now minimised by thickening these parts, and blanketing 
the outside of the roof with sand, keeping the construction together 
by very heavy bracing. 

B. iii. — Prevention of Cooling of the Hearth on Withdrawal and on 
Charging. — By far the most important cause of heat losses in working 
was occasioned by the withdrawal of the whole of the melted products, 
the charging of fresh cold ores, and the efficiency of the furnace, 
was very greatly reduced in consequence. In the older methods, fully 
three-quarters of the time and fuel, and almost all the labour, were 


spent in manipulating the charges and bringing them up to the point 
of fusion, the actual smelting operation being responsible for but a 
small proportion. The withdrawal of the hot slag and matte abstracts 
much of the heat of the furnace, and the cold charge which is fed 
in, not only cools the furnace hearth on which it rests, but being 
a poor conductor, prevents the heat from again penetrating through 
it to the hearth and to the undermost portion of the charge. It has 
been estimated through the use of pyrometers, that the temperature 
in the furnace after such withdrawal and recharging may drop to 
less than 700° C. — a dull red heat — and there is no way under 
such circumstances of heating up the hearth again, except by 
conduction through the charge. Some hours' hard firing were thus 
required to bring the furnace to the desired temperature again, after 
which it was necessary to re-open the working doors, in order to stir 
the materials so as to prevent the haK-fused masses, still lying on the 
hearth, from sticking to it. This also occasioned delay in the operations, 
and caused much waste of fuel, heat, and labour. 

B. iv. — Utilising the Heat of Melted Charges for the Heating of Fresh 
Additions. — All the above difficulties, and many others, have been 
overcome by maintaining a deep pool of hot molten matte in the 
furnace, and by feeding hot charges upon this matte layer. These are 
two of the most vital and successful changes introduced into modern 
reverberatory practice, and will be reviewed in detail subsequently. 

B. V. — Utilising the Heat of the Escaping Gases as much as possible, 
— ^Improvements in this direction have been brought about — 

(a) By constructing the furnace of as great a length as will allow 
of maintaining the charge in a sufficiently fluid state to permit of its 
being tapped from the furthermost end of the furnace. 

(6) By using the still hot escaping gases under boilers. 

Modern Reverberatory Practice, — The requirements for the 
; successful operation of the reverberatory furnace, and the methods 
for ensuring its efficient working which have just been reviewed, 
involve the application of the following principles, which are the 
essential factors in modern reverberatory smelting practice : — 

1. The grade of the furnace products is controlled at the roasters. 

2. The melting must be as rapid as possible. 

3. The employment of very large furnaces. 

4. The use of a heated matte-pool in the furnace. 

5. The charging of hot calcines. 

6. The regulation of the furnace working by draft pressures. 

7. The continuous working of the furnace. 

8. Modified constructional details. 


1. Control of Furnace Products at the Roasters. — This feature 
has already been indicated in deahng with roasting practice. The 
importance of this system in the economy and efficiency of the 
furnace working is very marked. 

(a) The roasting plant affords the most ready means of control 
over the desired sulphur elimination, this being its sole function. 
The modern roaster is so designed as to allow of almost perfect 
regulation in this respect, since amount of feed and rate of passage 
of the sulphides through the furnace are under perfect control. 

(6) The work of the reverberatory is thus confined to one object 
only, that of rapid melting down, to which the foreman can give his 
sole attention free from the necessity of manipulating the grade of 
the matte at the same time. 

In modern work it is usual to pass the whole of the charge 
(concentrates as well as flux) intended for the reverberatories, through 
the roasting plant. The advantages of such procedure are — 

(i.) The flux is preheated at little extra expense, there being usually 
plenty of heat to spare for this, and the roaster capacity is not unduly 

(ii.) Intimate mixing of the charge is assured, and this greatly 
facilitates the fusion and reaction. 

(iii.) More rapid and thorough roasting is effected, since the presence 
of the inert flux prevents clotting or undue sintering of the sulphides 
in the roaster. 

(iv.) The charge is found to be in a much better condition, both 
physically and chemically, for successful reverberatory smelting. 

Lime in the roaster charge appears to assist the thoroughness of 
the roast, whilst an incipient slag formation is commenced owing 
to the juxtaposition of basic oxides and silica, in the hotter parts 
of the roaster furnace. . 

2. Bapidity of meUing is an indispensable feature of modem 
work. The conditions necessary for rapid melting have been reviewed 

3. Use of Large Furnaces. — Reverberatory furnaces appear to have 
replaced the blast furnace in Great Britain somewhere about 1700, 
and by 1854 they were in general use in this country. At this period 
the usual dimensions were, for the hearth 13 feet by 9 feet, with a 
fire-box 4 feet by 4 feet, the furnace having a capacity of 12 tons per 
twenty-four hours. In Great Britain the size increased very slowly, 
and it was in the United States of America that the important 
increase in dimensions and in enormous outputs were developed. 



The work was commenced systematically in about 1878 by Richard 
Pearse (a Swansea-trained metaUurgist) at the Argo Smelter in 
•Colorado. Table V. indicates the gradual improvements in practice 
resulting from these developments (see also Fig. 23, p. 90). ' 

Table V. — Developmknt in Size of tup: Reverberatory Furnace. 






Tons Ore per 
Ton Coal. 

1878, . 

4' 6" X 5' 

9' 8" X 15' 

2' 9" 

12 tons. 

2-4 tons. 

1882, . 

4' 6" X 5' 

10' 4" X 17' 10" 

2' 9" 

17 „ 

2-43 „ 

1887, . 

4' 6' X 5' 6" 

12' 8" X 21' r 


24 „ 

2-67 „ 

1891, . 

4' 6" X 6' 

14' 2" X 24' 4" 


28 „ 

2-8 „ 

1893, . 

5' X 6' 6" 

16' X 30' 

3' 6" 

35 ,,(43)* 

2-7 ,,(3-3)* 

1894, . 

5' X 6' 6" 

16' X 35' 




1903, . 

5' 6" X 10' 

20' X 50' 

5' 5" 



1910, . 

8' X 16' 

19' xll6' 



This practice has been continued in modern smelter work, the 
developments being in the direction of attempting to melt the largest 
possible quantity of charge in one furnace as rapidly as possible. 
This has been found to depend upon the rapidity with which the 
<uel is burned, and the enlarging of the fire-box had a, specially im- 
portant influence in effecting this rapidity of combustion. 

Then, with the size of grate fixed and the most efficient burning 
i the fuel arranged for, the capacity of the furnace depends simply 
_ on increasing the area of the hearth to as great an extent as the heat 
generated is capable of maintaining at the desired temperature. 

The breadth of the furnace is however, limited by — 

(a) The span of arch which can be supported in the construction. 
(6) The length of the tools which can be conveniently managed. 

The maximum width so far found satisfactory is about 19 feet, so 
that this dimension being fixed, the furnace capacity is enlarged b^ 
increasing the length, and this is limited only by the distance from 
^he fire-box to which the flame can maintain the temperature necessary 
or keeping the charge in a state of perfect fluidity. For many years 
he length was regarded as limited to 50 feet, smelting about 2-7 to 
i tons of charge per ton of coal, but E. P. Mathewson, at Anaconda, 
finding the escaping gases still very hot, gradually increased the length 
of the hearth, first to 60 feet, then to 80 feet, and finally up to 116 feet, 

♦ The charges of calcines were fed whilst .still rod hot. 



Fig. 23. — Development of the Reverberatory Furnace (Gowland). 

when the furnace smelted 4-83 to 5-0 tons of charge per ton of coal. 
The gases then left the furnace at a temperature of about 950° C, 
and contained sufficient heat to fire two Stirling boilers, each of 



375 H.P. Every furnace thus provided about 600 H.P. from this 
^ waste heat, and the gases finally escaped at a temperature of 320° C. 
The capacity of these large furnaces is about 270 to 300 tons of 
charge per day, and in addition to the economy and efficiency resulting 
from the treatment of such large quantities of material at once, there 
are the further great advantages in that — 

(a) Setthng of matte and slag is much more perfect when such 
large quantities of fluid material are stored. 

(b) Tapping of matte and slag is easier and more efficiently conducted. 
About 110 feet appears to be the practicable maximum for furnace 

length, and reverberatories of this size are being constructed wher- 
ever circumstances permit, several new smelters having erected such 
furnaces — there are eight at Anaconda, Mont. ; two at Garfield, Utah ; 
five at Tooele, Utah ; four at Cananea, etc. The length of the hearth 
is naturally dependent upon the character of the fuel, particularly the 
length of flame given out on burning. Bituminous fat coals are the 
most suitable for this purpose, and in localities where such fuel is not 
available, the use of liquid fuel has now been successfully adopted. 

4. Maintaining a Heated Matte Pool in the Furnace. — This is- 
probably the most important and beneficial advance made in rever- 
beratory practice. 

In certain stages of the old Welsh process, a store of matte was- 
retained in the furnace after skimming off the slag, but the object was 
to collect a sufficiently large quantity of matte in the furnace for 
convenient tapping out. 

The modem practice has several objects and possesses enormous 
advantages — 

(i.) It assists efficient settling. 

(ii.) It conserves the heat inside the furnace. 

(iii ) It presents a highly heated surface for the fresh charge 
to fall upon, and thus greatly increases the rapidity of melting, by 
ensuring that the charge is heated both from above and from below. 

(iv.) It prevents the sticking of half-fused charges to the furnace 
bottom, the removal of which masses would necessitate much labour, 
and occasion cooling of the furnace by the opening of working doors. 

(v.) It preserves the furnace bottom. 

Liquid matte has practically no action on the siliceous material 
of the hearth, and so presents an inert mass between the bottom and 
the charge. This charge consists of calcines (mainly oxides of iron), 
which would, during the process of melting down, slag with and corrode 
he furnace hearth were it not protected by the matte layer. 





(vi.) It allows of continuous charging and withdrawal of materials, 
and of continued high temperature in the furnace, thus protecting 
the furnace lining from much wear and tear. Nothing damages furnace 
linings more than exposure to changes of temperature, on account of 
the continual expansion and contraction of the brickwork and the 
low thermal conductivity of the silica. Furnace linings wear out much 
more from such action than from long exposure to continued high 

(vii.) There is effected an enormous saving of time, fuel, and 
labour by maintaining a constant high temperature, instead of having 
to heat the furnace up again after each tapping and charging, as was 
the case with the older methods of working. 

(viii.) The levelling of the charges in the furnace is greatly facili- 
tated. The charges would otherwise pile up under the charging hoppers, 
and form heaps which are not only difficult to melt down, but which 
tend to stick to the furnace bottom, requiring time and arduous labour 
for their removal. In modern practice, charges in quantities of 10 to 
15 tons at a time may be dropped in, these merely spread themselves 
out on the bath of molten material and float down in a thin stream 
towards the skimming door at the end, and they generally melt and 
disappear when haK-way down the furnace. 

By this means, the working doors at the side need practically 
never be opened for manipulating the fresh charges. 

5. The Charging of Hot Calcines. — This improvement was also 
introduced by Pearse, and possesses very many advantages ; he was 
able to increase the furnace output by 23 per cent, with the aid of this 

Instead of allowing the materials from the roasters to cool down, 
they are taken straight from the roaster bins to the hoppers which 
feed the reverberatory furnace, where they retain much of their heat 
until charged into the furnace, being then still red hot as a rule. 
Much time and fuel is thus saved owing to the charge requiring less 
heating up, and the cooling action of charging is diminished. 

A charge of 15 tons is completely melted within an hour. 

6. Regulation of Furnace by Draft Pressure. — It has already been 
pointed out that rapid combustion of fuel, and consequently rapid 
melting, is greatly assisted by good draft through the furnace. In 
modern practice, where the factors, such as charge composition, nature 
of fuel, and furnace proportions, have been satisfactorily arranged for 
independently, the actual working of the furnace is regulated by 
the draft pressures. These are registered automatically by water- 
manometers arranged at various points. One usually communicates 




mth the furnace, above the fire-bridge ; another is connected to the down- 
take flues. The indications of these instruments enable a record to be 
kept of the various operations, and of the charging of the furnace, as well 
as of the condition of the fire. The usual draft pressure worked with 
corresponds to about 0-8 inch of water, registered above the fire-bridge. 

On opening the hopper for charging, the pressure drops almost to 
zero ; the opening of any doors causes a reduction in pressure ; the 
charging of coal is also rendered noticeable by a drop in the record. 
Reduction of pressure also indicates " airing '" of the furnace by an 
excess of air entering through channels in the bed of coal ; draft- 
pressure thus acting as a check on the firing and also on the grating, 
since the formation of excessive clinker in the fire-box is indicated 
by an increase in the pressure. 

Corresponding to such record over an 8-hour shift, as shown on fig» 
24, Offerhaus noted the following furnace manipulations, illustrating 
how accurately the operations are checked by this method : — 


7.00-7.14 Skimming (coal charged during this period). 

7.16-7.16i Side door opened. 

7.28-7.31 Coal charged. 

7.52-7.57 Charged. 

8.05-8.15 Tapped. 

8.15 Coal charged. 

8.40 Coal charged. 

8.54-8.59 Grating. 

9.05 Side door opened. Charged. 

9.27 Coal charged. 

9.49 Coal charged. 

10.07 Charged. 

10.25 Coal charged. 

10.41 Coal charged. 

10.45-10.58 Skimming. 

11.04 Coal charged. 

11.16 Charged. 

11.16-11.35 Some grating. 

11.36 Coal charged. 

12.03 p.m. Coal charged. 

12.04 Charged. 

12.37-12.481 Tapped, \\ ladles (about 11 tons). 

12.45 Coal charged, 

1.00 Charged. 

1.11-1.45 Grating. 

1.26 Coal charged. 

1.44 Charged. 

1.51 Coal charged. 

2.18 Coal charged. 

Total charges during shift, 

16 coal, 7 calcines. 



The draft record is placed close to the chargmg platform, in order 
to be in a convenient position for the guidance of the workmen. The 
draft in the main flues is 1-7 to 1-8 inches water pressure ; this is 
similarly recorded in the foreman's office. 

7. Continuous Working of the Furnace. — The continuous working 
of the furnace is a most important factor in modern practice, and is 
naturally inseparably bound up with the principle of maintaining the 
heated matte-pool in the furnace, which allows of the continuous 

Fig. 24. — Draft Pressure Record of Anaconda Reverberatory Furnace (Offerhaus). 

-charging of hot " calcines,'' and the continuous or regular withdrawal 
of slag and of matte when required. 

The matte (which can be efficiently settled, owing to the prevailing 
high temperature and the large mass of heated material in the furnace) 
is stored there until required at the converters, when the desired 
quantities are tapped out. The slag which is produced by the smelting 
action gradually accumulates, and at regular intervals most of it is 
run out (rather than skimmed). This usually takes place every four 
hours. The slag accumulates until it reaches a level some 3 or 4 inches 
above the skimming plate at the end of the furnace, and the quantity 





which is run out at each " skimming '' amounts to some 60 or 80 tons, 
the contents of the furnace being lowered to such an extent that a 
fresh accumulation of material may proceed during the next four hours. 
No pulling of the slag is required as in the older methods of working, 
since the material is so very hot and fluid that it simply pours out 
of the furnace, and twenty minutes usually suffices for the whole of 
the 60 or 80 tons to run off, the rabble being used chiefly to regulate 
and control the stream, and to keep back siliceous crusts or floaters. 
The slag is run out until the matte is seen underneath, on flapping back 
a thin layer, or until the level of the skimming plate is reached, and 
its removal is such a short and simple operation that there is very 
little interference with the regular and continuous running of the 
furnace. Similarly, the tapping of as much as 50 to 100 tons of matte 
from the store of 250 tons of hot fluid material has little influence on 
the continuous working. Charging of coal and calcines is performed 
at regular intervals, and the charges of 15 tons of " calcines '' fed in at 
a time, readily melt down and settle. Practically the only interference 
with continuous running is the necessity for claying and repairing, 
and the use of the matte pool on the hearth has lessened the frequency 
for this to a large extent, the hearth bottom itself being protected 
from corrosion, owing to the sulphides exerting no action upon it, 
whilst the oxides in the charge which would be capable of attacking 
the siUceous bottom are slagged off before they get an opportunity 
of reaching it. The hearth bottom, if properly put in, is practically 

The portion of the furnace most subject to corrosion is at the slag 
line, where deep channels are gradually cut out. Every four to six 
weeks the furnace is tapped dry, repaired, and fettled, as much as 
20 tons of fettling sand being often required for this purpose. The 
sand is thrown in and patted into place by long rabbles, the operations 
occupying about eighteen hours. Every nine months or so the furnace 
is repaired more fully, 20 or 30 feet of brickwork near the fire-bridge 
being taken down, and the great cavities in the side walls repaired by 
masons, using sihca bricks. The employment of higher temperatures 
I in modem work allows of more siliceous slags being produced, which 
lessens the tendency to the eating away of the walls. 

The feeding of siliceous copper ores through a series of small 
lioppers situated in the roof, near to the walls, has lately been intro- 
duced with a view to protecting the furnace sides from the corrosive 
action of the slag, and to exposing a suitable siliceous flux to this 
jnaterial. This appears to have fulfilled its purpose to some extent, 

t," — 



the tendency for the cold added material to form floaters, which 
require limestone additions in order that they may be fluxed off ; and 
the cooling effects and leakages through the openings have also given 

8. Modified Constructional Details. — In addition to the increased 
size of fire-box, hearth, and flues, and to the necessity for very 
heavy staying in order to keep the enormous arch in permanent 
shape, which are characteristic of modern practice, the construction 
of modern furnaces involves the building of a suitable hearth to 
carry the heavy burden of hot and fluid matte which is stored in the 

It was formerly considered correct practice, in the smaller types 
of furnace, to construct the hearth over a vault, in order to keep the 
underside cool and thus prevent the corrosion and eating away of the 
sihceous bottom by the oxidised charges, during the process of melting 
down. In modern practice it is absolutely essential to work with a 
perfectly solid structure. 


Fig. 26. — Transverse Section of Modern Reverberatory Furnace, Anaconda, 
indicating Foundations, Hearth, and Bracing. 

(a) Because the hearth must be kept as hot as possible, so a» 
to ensure rapid melting of the charge and maintain the products in 
a perfectly fluid condition. Any circumstance tending to cool the 
hearth is rigorously avoided, this being the contrary of the older 
practice. The protective influence of the heated matte-pool in modern 
work preserves the bed from the corroding effects of fresh oxidised 
charges, and in consequence, the maximum degree of heat can with 
safety be maintained on the furnace hearth. 

(b) The enormous weight of charge and the heavy arch and walls 
demand the strongest possible foundations and support. 

In building modern reverberatories, the foundation for the hearth 
is constructed of solid masonry or brickwork, or as at Anaconda, of 
a solid bed of slag, some 24 inches in depth, run in from an adjacent 






w'^Kml Z^S^^Ini^^^^' 





furnace. The I-beams used for carrying the bracing are erected in a 
surrounding trench, and a further quantity of slag^(4 feet thick by 
2 feet deep) is run in, thus yielding a perfectly rigid and impervious 
foundation (Fig. 26). On the top of this slag-foundation is built 
a layer, 12 inches thick, of sihca bricks, and upon this, the actual 
worknig bottom of the furnace is constructed. 

This bottom is now put in also in a manner different to the older 
practice, and excellent results have accrued from the change. 

The old method of constructing sand bottoms consisted of putting 
in the beds of sand, layer by layer, and thoroughly fritting each one 
before the addition of the next ; in modern practice, it is found that 
proper consolidation is not attained with beds of the enormous 
area now employed, when the bottom is constructed in such 

The present method of working the reverberatory furnace is not 
to drop the charge on to the sand hearth at all, but into the deep 

j pool of matte, and the sand-hearth is regarded more as a convenient 
foundation for the support of this liquid working-bed, on account of 
Its constituting a cheap non-conducting and fire-proof material which 

I is unafiFected by the materials resting upon it. It was found, however, 
on commencing this matte-pool practice, that the older method of 
putting in the bottom in successive sand layers was not suitable for 
this work ; after a Httle wear, the beds became raised in layers, this 
l)eing especially the case if any holes happened to be eaten through 
in places. Moreover, the large weight of matte tended to find its 
way down between the layers and raise them up bodily, or else it 
worked down at the edges of the hearth and side walls, and either 
broke out underneath the former or through the latter. When it 
was ascertained that Hquid matte itself had no corrosive action 
on the siliceous hearth if the latter be kept constantly covered, and 
that the causes of breakouts were principally due to mechanical weak- 
nesses, it required only improvements in design and construction in 
order to avoid them. This is now attained by constructing the bed in 
a compact and perfectly massive form, and is best accomplished bv 
'Utting in the whole layer of 26 inches of sand at once, and firing as 
iiard as it is possible for the brickwork to stand. The method has 
met with exceptional success in practice, rigid and impervious hearths 
are obtained ; it being found that less than 1 inch has worn ofi^ the 
^»ed after two years' working. 

K Large Reverberatory Furnaces: Details of Construction,— 
e large furnaces at Anaconda were the first of the modern type to 
constructed, they have met with enormous success in practice 



and constitute the standard form. Similar furnaces are now in opera- 
tion or under construction at many of the large modern camps, and 
are of similar design and construction. 

The hmrth is 102 to 116 feet long by 19 feet wide. 

Grate, 16 feet by 8 feet = 128 square feet grate area. 

Ratio of hearth to grate area is 16 : 1. 

Distance from hearth to level of fire-bridge, 26 inches ; hearth to 
crown of arch, 6 feet 5 inches. Walls are 26 inches thick. Roof is 
15 inches thick (except for 4 to 6 feet over the fire-bridge, where it 
is 20 inches). The bracing of the furnace is necessarily particularly 
strong (see Fig. 29). Lined inside with silica brick, said to be the 
finest in the world. The bed is of the finest Dillon sand (97-5 per 
cent, silica), ground to pass J-inch mesh ; the bed has a slope of 
8 inches towards the tap-holes, of which there are two. During the 
construction of the large furnace there are left in the roof ten 
expansion openings of 3 inches each, which by the time the furnace 
has attained its working temperature, become closed up (see Fig. 30). 
The conker plate which runs through the fire-bridge is 14 to 15 feet 
long, and is made thicker near the furnace side, where it is 3 inches 
thick. The air space through the plate is 2 feet 3 inches by 9 inches, 
and serves the purpose of keeping the fire-bridge cool ; air passes 
through it continuously, and if the plate shows signs of becoming 
hot, a blast of cold high-pressure air is sent through it. Still 
further heating of the plate and signs of red heat are an indication 
that the 2 feet of silica of the fire-bridge wall are being burnt 

Working of the Revcrberatory Plant at Anaconda. — The 
plant consists of eight large furnaces, built parallel to one another, 
seven being usually at work whilst the eighth is undergoing repair. 
Each furnace treats 300 tons of hot calcines and flue-dust daily. 

Charging. — The furnaces are charged every 65 to 70 minutes with 

15-ton charges, and as soon as one charge is melted, another is added ; 

with average running, 150 charges are worked in the seven furnaces 

daily. The charge train, consisting of an engine and three cars, each 

of which ca,rries 5 tons of charge, travels from the roasters and enter."^ 

) the revcrberatory building by an overhead track running above the 

^harge bins of the furnaces. It discharges through hoppers into the 

ins which extend across the entire width of the hearth. Bins were 

rmerly arranged at intervals all the way down the furnace, but now 

"uly the two bins nearest to the fire-bridge are employed. Into the 

l>'ack bin, 10 tons of charge are placed, and into the other, 5 tons. 

f Each of these bins discharges through two hopper discharge openings, 


feeding the furnace through holes in the root (Figs 29 30, which 
are closed when not in use, by round firebrick tiles 20 mches in 
dlLeter a„d 21 inches thick ; these are moved in and out of position 
bv means of lev'ers operated from the fire-box platform. 
' The temperature maintained in the furnace is high, approximating 
to 1 500° C and iust previous to dropping in a fresh charge, a work- 
man'Ty means of"a rabble, feels about the hearth below the charging 
hopper in order to ensure that all of the previous charge has been 
meHed, and that none of it is sticking to the "rnace l^arUi. By 
employing only the comparatively small quantities of 15 ons, thm 
sticking if avoided, since such charges are not heavy enough to sink 
™Ued through the 8 inches of slag and 8 mches of matte in 
the furnace. The former practice of feeding charges amounting to 
45 tons through hoppers situated all the way along the tenace had, 
given serious trouble in that respect, and had consequently to be, 
discarded. When the examination of the hearth is completed, he 
time occupied being very short, the side door is closed, and sealed^ 
Sh sand the covers to the holes in the roof are now withdrawn, 
the gates closing the hoppers pulled back, and first the ^-ton theni 
the fo-ton charge is dropped into the furnao^. The whole operaj 
tion, including the preliminary opening of the door to test the furnaceJ 
bottom, occupies five minutes. 

Very Httle hand labour is required round these enormous furnaces, 
except for the grating of the fires, for the charging of coal and calcines 
every hour by the operation of levers from the fire-box platiorm, 
for the skimmfng of slag at intervals of four hours, and for the tapping 
of matte when required. The whole of this work is conducted by the 
skimmer and two helpers to each furnace, one of the men also looking 

after the boilers. .it u^„ 

As soon a« the charge has been dropped on to the pool of molten 
material, the mass appears to spread out over tiie surface and float 
towards the skimming door, in a thin slow-moving stream which 
disappears when about half-way down, being usually melted withm 
one hour. The former 40-ton charges required as much as eight hours 

**"' Owh" to the great heating effect of the large bath of hot material 
below, and of the intense flam^e above, there is but Uttle coohng action 
on adding the fresh charge ; whilst with this length of furnace practi- 
cally all the dust is settled, and very httle is carried into th^ Aues. 

CoaUng.-The quantity of coal employed amounts to 20 to 2o per 
cent, of the charge, or about 50 to 60 tons per day per furnace, 1 ton 
of coal smelting rather less than 5 tons of calcines. 

[To face page 100. 

Fig. 29. — Fire-box End of Reverberatory Furnace, showing massive Bracing, Charge Bins, 
and Charging Levers — Anaconda. 

Fig. 3(). — Interior of Reverberatory Furnace (looking towards Skimming Door), 

aVi/iiir;n» V^-..^..-.;^ — a :„ T)„..f i rn 

T1T,.1„_ A, 


Coal is charged every 40 minutes in quantities of 1 1 tons at a time, 
from bins which extend across the entire width of the fire-place, feeding 
through four hoppers into openings 1 foot square in the roof of the fire- 
box, and the withdrawing of the gates is operated by means of levers 
at the platform. Over the fire-bridge are two rows of air-holes used 
for regulating the length and character of the flame in the furnace ; 
the flame, however, plays a subordinate part in the smelting reactions. 
The coal employed is from Diamonds ville, Wyoming, and gives a flame 
125 feet in length, the appearance of which is gauged through the 
window fixed in the off -take flue, this being visible from the fire-box 
platform. The coal is run-of-mine quality, and considerable slack is 
used. It possesses a high calorific power and a large proportion of 
volatile constituents, but clinkers rather badly, and a clinker grate is 
worked with. 

Grating. — The fire rests upon 3-inch round bars placed at 4| to 
6-inch centres, and is maintained at a depth of about 27 inches. 
Gratmg requires to be conducted at fairly frequent intervals, usually 
twice per shift, in order to keep the fire free and to prevent channelling, 
which is indicated on the draft gauge by a drop from 0-75 inch to 0-50 
inch, due to airing. It serves further to prevent clinkering, which, 
when taking place in the fire, causes a rise of from 0-75 up to 1-0 inch 
on the gauge. The operation of grating usually occupies about half- 
an-hour ; the work is arduous, and the heat to which the workman 
is exposed is itself very trying. 

Coke Recovery. — A constant stream of half-burnt fuel and ashes 
falls through the bars, and during the clinkering operations large 

juantities are dropped. The material all falls down a bank inclined 

a 45°, into a channel where it is met by a stream of water which washes 
it along launders and through a grizzle, to a settling tank. The settled 
products are subsequently jigged, the recovered coke being washed 
over the tail-board to a trommel, and by this means 10 per cent, of 

he fuel charged into the furnace is recovered in a useful form. This 

oke is used up as a constituent of the briquettes. 

Tapping the Furnace. — Matte is usually withdrawn from these 
large stores upon such occasions as it is required for the converters, 
though sometimes when the supply has got ahead of the converters' 
^mands, the matte is tapped and run outside the reverberatory 

(uilding, being cast into large matte-beds. The tap-holes are situated 
stween the second and third doors, and between the fourth and fifth ; 
id each consists essentially of a copper plate 2 inches thick and 

f6 inches square, which at first stands back 9 inches from the 
itside of the wall. Through this plate a 1-inch hole has been 



Ladles of 
Matte in 
at End of 

O O O O O O 

,—1 ^H rH r— 1 I— ( r— 1 

; O 


Draft, 1-7 inches. Number of furnaces running, . . . 7-00 

All furnaces working slow. Number of charges, 140 

Furnace No. 5, one bad charge. Ladles matte tapped, .... 34 

Cupriferous material smelted per furnace, 275-6 tons. 












(M X --O :o CO 00 

: ^ 


Draft, 1-7 mches. Number of furnaces running, .... 6-67 
Furnace No. 1 delayed 8 hours tapping and claying. Number of charges, . . . . .118 
Furnace No. 7 down for repairs. Ladles matte tapped, . 47 
Bad coal on all furnaces. Cupriferous material smelted per furnace,' .* 24(5-7 tons. 

S Tt< Tl^ tJ< -* -* -# 
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Cost of 

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drilled. The tapping bar is maintained inserted up this hole, being 
passed through the conical clay plug which closes it. At the back of 
the plate is 21 inches of lining material through which the tapping- 
hole passes. When the copper plate shows signs of a red heat, it 
is an indication of the lining tending to burn through ; this part 
of the furnace is then cooled, the plate taken out, a 9-inch layer 
of sand is rammed into position, and the plate is thus moved forward 
a corresponding distance. Such a tap-hole plate lasts for about five 

The reverberatories are usually not tapped until they contain 
about 250 tons of matte. The operation of tapping is performed by 
withdrawing the rod by means of a wedge and ring, when the matte 
flows along the launders leading to the ladles for the converters ; two 
ladles of about 8 tons capacity each are usually filled at once, each 
ladleful being sampled at the runner. The tap-hole is then stopped 
with a cone of clay, and the tapping-rod driven through it again. 

Typical daily reports of the furnaces are appended in Tables VI. 
and VII., and a monthly report on Table VIII. 

TABLE VII. — From Daily Assay Report — ^Reverberatory Furnaces. 

August 19, 1908. 

Per Cent Copper in Slag. 

Furnace Number. 

Shift 1. 

Siiift 2. 

Shift 3. 






























Average in slag, 





rSiOj, 29-5] 

per cent. 

mposition of 

FeO, 37-3 


Composition of slag 


S, 7-7 
CaO, 2-7 



Copper, 8-« 


Copper in matte, . 

{SiOg, 39-4 per cent. 
FeO, 40-7 „ 
CaO, 4-3 „ 


Fuels for Reverberatory Furnace Work. — The chief require- 
ments of the fuel for good reverberatory work will now be apparent., 
particularly with regard to length of flame. This depends to a large 

















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extent upon the proportion of volatile hydrocarbons, but also on the 
conditions under which they are given off. For instance, a coal which 
rapidly parts with its hydrocarbons and leaves in the grate a dense 
layer of slow-burning coke would be unsuitable for reverberatory 
work, though some caking is necessary in order that the fuel should 
not burn away too rapidly, as it should yield a good bed of the 
required depth. 

The great success of large reverberatory furnaces worked under 
suitable conditions, has had the tendency to tempt smelters in different 
parts of the world to erect furnaces of similar size independently of 
the character of the available fuel, and in several cases results have 
been unsatisfactory, at least in the earlier stages. 

These preliminary failures have, however, served the purpose of 
developing the adaptation of other fuels for this work, and from the 
employment of oil for the purpose, important extensions in practice 
will undoubtedly develop in the future of reverberatory furnace 

The device of using pulverised coal as a fuel has attracted atten- 
tion at several smelters where the local coal as mined was proved to 
be unsuitable for use. In practice, however, the method has, up to 
the present, given unsatisfactory results, for although a longer flame 
and higher temperature have been obtained in the furnace, difficulties 
in working have arisen which appear to bar its use. One of the 
chief drawbacks has been due to the fine ash from the fuel, which is 
deposited in the flues in large quantities and even causes considerable 
slagging in them, impeding the working of the furnace and pre- 
venting the recovery of heat from the furnace gases. Further difficulty, 
though not quite so serious, was caused by the dust being blown upon 
the charge and tending to settle upon it, forming a non-conducting 
I blanket which retarded the melting of the material by the flames. 
The method does not appear at present to offer much promise of 
X tended application to copper smelting. 

Oil Fuel in Reverheraton/ Practice. — The successful application of 
•il as a fuel marks a useful advance in reverberatory practice, parti- 
ularly in connection with the working of large furnaces. 

On several of the smaller plants, oil fuel has been in use with 
onsiderable success for some time, but within recent years the 
building of large-sized furnaces without having at hand suitable coal 
resources has led to attempts to employ oil in its place, and the pre- 
liminary difficulties appear to have been to a large extent successfully 
overcome. The work at the Cananea Smelter with oil fuel, and the 



indications of the possibilities of this method. Working on char^ 
consisting to a large extent of flue-dust, several thousand tons 
material have been smelted in furnaces yielding 245 tons daily output, 
at a cost which compares very favourably with that of ordinary 
practice. This success is particular^ noteworthy in view of certain 
features in the preliminary system of working which will doubtless be 
altered at no very distant date, and of the fact that flue-dust is 
sometimes a difficult material to melt in a reverberatory furnace 
even when good coal is available as a fuel. 

6 inches 

r-^Pipe Ff^nof \ 

Fig. 31. — Shelby Oil-burner for 
Eeverberatory Furnace Use. 

The chief difficulties in working appear to have been largely in 
connection with the regulation of the flame and the management of 
the oil-burners. In endeavouring to obtain the requisite high tem- 
perature over the entire length of the furnace-hearth, an intense local 
action was caused near the place where the oil in the form of a spray 
entered the furnace, resulting in the burnino: out of the roof-arch on 
several occasions. These difficulties will doubtless be overcome with 



further experience in the design and management of the burners 
constructed for this class of work. 

At Cananea, four oil burners of the Shelby type are employed on 
each furnace, and this form is stated to project the flame further into 
the furnace, and to prevent its impinging on the roof, more successfully 
than the other types tried. The waste heat fires three Stirling boilers 
of 664 H.P. Less than one barrel (42 gallons, or 310 Ibs.^l of oil is 
consumed per dry ton of charge, and of this quantity 0-43 barrel 
is chargeable to steam-raising under the boilers. The manner 
of working the charges, and the furnace construction in other 
respects, follow very closely the methods of operation already 

Costs of Oil-fired Reverberatory Working* — Ricketts has con- 
tributed a useful analysis of the costs of reverberatory work using 
oil as fuel, under the conditions prevailing at Cananea, Mexico. He 
noted that the use of too much oil should be avoided. This precaution 
led to a decrease in the amount of repairs necessary. 550 barrels of 
oil were required to get the furnace into fairly good condition, and 
8 barrels per furnace per hour to keep it going well. It is hoped 
ultimately to reduce the oil consumption to 0-8 barrel gross per ton 
of charge. 

Analysis of Oil-fired Reverberatory Furnace Costs— Cananea— February 

Tonnage Chabged — 

Calcines, . 


Operating expenses, 
Slag and matte expense, 
Boiler-house, . 
General expense. 
Cost of Hux, 

Steam credit, . 
Operating cost, 

to July, 

1911, inclusive. 


5 Days, 312-5. 

Dry Tons. 

Per cent, of Total 


. 21,019 
. 35,533 
. 3,040 









Per Dry Ton. 

nse, . 

. $111,687-17 





SI -8593 







Analysis of Costs — 
(1) Operating — 

Labour, . 


Fuel oil, . 







Per Dry Ton 






1 -4654 













(2) Repairs — 
Labour, . 
Supplies, . 
Shop expense, . 

. $11,063-93 







Steam credit, . 

. $133,303-05 


. $84,441-19 


Net total, . 


Gaseous Fuel. — The proposal to employ gaseous fuel in copper 
smelting dates from the introduction of this method of furnace-firing 
by Siemens 50 years ago. It is, however, not in general use, although 
at several smelters gas-firing is employed in furnaces for the refining 
of the metal. 

The chief difficulties have been in connection with the control of 
the flame, burning-out of the roof having been a not infrequent occur- 
rence when employing gaseous fuel, and the method has been tried 
and given up at the Great Falls Smelter in Montana, and at several 
other works. 

The practical difficulties ought not, however, to be insuperable 
should gas-firing be otherwise found most practicable for the parti- 
cular conditions at the smelter, although there appear to be certain 
physical characteristics of such flames which may be responsible for 
some of the difficulties met with in employing this type of fuel for the 
working of very large reverberatory furnaces. 

The Condition of the Charge for Good Reverberatory Work, 
— The considerations which decide the advisability or otherwise of 
installing at a smelter, any particular types of furnace, whether rever- 
beratory or blast furnace or both, cover a very wide field, and will be 
more apparent when blast-furnace practice has been reviewed in detail. 
It is clear that the blast furnace is unsuited for the direct smelting 
of fine materials as such, and that the reverberatory form of furnace 


is best fitted for their treatment when large quantities of this material 
require to be dealt with. Actual practice has shown, however, that 
the reverberatory does not give equally satisfactory results on all 
classes of fines, and that there are certain physical and chemical 
conditions of the charge which appear to be necessary for the most 
successful and rapid smelting. When such conditions are not adhered 
to, less satisfactory working has resulted. Recent experience has, to 
some extent, defined more clearly the nature of these requirements, 
and has indicated the procedure which is necessary in order to avoid 
an undue supply of the less suitable material for the reverberatory 

It is usual to smelt in the reverberator}^ furnaces, where such are 
available, the greater portion of the dust which accumulates in ^ery 
large quantities in the flues at the smelter. The reverberatory is 
the only type of furnace in which such material could be treated 
directly, under the present conditions of working. In practice, 
however, it has been found in several instances, though not 
universally, that such dust is considerably more difficult to treat 
in the furnace, and entails considerably more expense in smelting 
than does the ordinary roasted concentrate. It is estimated by 
Ricketts that this extra cost is practically equivalent to the expense 
of roasting an equal weight of concentrate. 

Flue-dust, as a rule, consists mainly of material in a minute state 
of division, in which condition, as is well known, a much higher tem- 
perature is required for its fusion than if it were in the form of 
coarser particles. This is largely due to the poor conductivity for 
heat which generally characterises such dust, and to the insulation 
by the air envelopes surrounding the individual grains, which thus 
prevents the heat passing from particle to particle, and retards 
their clotting, even when the prevaiUng temperature would otherwise 
be sufficient to cause fusion. The particles of flue-dust moreover. 
I *have been blown from the surface of the charge, especially in the blast- 
furnace process, and are thus rapidly and often almost completely 
oxidised in passing through the oxidising atmosphere which prevails 
ibove the charge and in the flues. Such oxides clot only with the 
greatest difficulty, and are characterised by comparative infusibility 
find poor conducting power, and hence are found to melt with con- 
iderable difficulty when treated in the reverberatory furnace.* 

Roasted fine concentrate, on the other hand, constitutes an ideal 
material for the reverberatory furnace charge, and the system of 

♦ The East Butte Copper Mining Company has recently reported the successful sintering 
'f its flue-dust by Dwight-Lloyd machines. (See Mining Journal, Jan. (5, 1912, p. 21). 


passing both the concentrate and the flux through the roasters hi 
been shown to possess numerous advantages. In addition to the 
thorough mixing and the preheating of the furnace charge, it was 
found that its chemical and physical conditions were particularly well 
suited for the subsequent reverberatory furnace treatment. The 
particles of concentrate, being gradually heated and constantly stirred 
in the presence of the small proportion of flux usually required, roast 
well, and lose the desired quantity of sulphur without an undue 
amount of preliminary clotting which w^ould otherAvise interfere Avith 
the operation, whilst any residual sulphide in the product is uniformly 
distributed through the roasted charge. In addition, at the higher 
temperatures which prevail in the later stages of the roasting process 
Avhen almost as much sulphur as was desired has been driven off, the 
materials are raised to a point approaching incipient fusion and slagging. 
The heat in the reA^erberatory furnace is sufficient to complete this 
effect, and enable the necessary chemical combinations and physical 
separations to be readily accomplished. 

The roasted concentrate should therefore form the main pro- 
portion of the reverberatory charge, Avorking in Avith it, in moderate 
quantities, such flue-dust as is made at the smelter. Of this flue- 
dust, it is naturally desirable to produce as small an amount as 
possible, not only on account of the difficulties in subsequent treat- 
ment, but also on account of the actual losses in the economj'^ of 
the furnace processes and the cost of rehandling, etc. In modem 
smelting, naturally, every effort is made to reduce the quantity of 
dust to the lowest practicable limit. 

The greater portion of the dust results from the treatment of 
unsuitably fine material in the blast furnace, and by decreasing the 
quantity of this constituent the flue-dust problem Avill be largely 
overcome. The smelting of fine concentrate in the blast furnace 
has up to the present been considered judicious where circumstances 
have rendered imperati\"e the addition of sulphides to the charge 
irrespective of their physical condition (either to act as a base for 
the matte, or on account of their fuel values), though naturally the 
proportion of fines has been kept as low as possible. 

The recent dcA^elopments in sintering processes, however, suggest 
the possibility of the future successful treatment, after preliminary 
agglomeration, of fine concentrate in the blast furnace, and if it be 
found possible to conduct the sintering by utilising the heat of 
oxidisation of the more free sulphur atom of the pyrites, and thus 
leave the bulk of the iron-sulphide fuel A^alues in the sintered product, 
as suggested by Peters, the difficulties in connection Avith excessiA^e 


flue-dust production from the above causes will be largely overcome, 
and the reverberatories will thus be relieved of this difficult constituent 
of their charge. 

It therefore appears desirable, when circumstances permit, either 
to agglomerate fine concentrates and then treat them in the 
blast furnace, or else to roast them and smelt the product in the 

So far as present experience has gone, it appears that — other 
circumstances being equally favourable — the correct scheme of 
treatment depends almost entirely upon the composition of the 
concentrate, there being for each process a particular class of 
fines for which it is best suited. The sintering process deals most 
satisfactorily with one class of concentrate, whilst the roasting 
process seems more particularly suited for a different type of 

Thus the higher the iron and sulphur values, and the lower the 
silica content, the more successful, cheap, and efficient is the roasting 
process — the Anaconda material for example roasts well, requires 
practically no external fuel or heating, and with the added flux, 
works very successfully in the reverberatories. 

As the silica content increases, however, and the iron and sulphur 
contents diminish, there is a consequent decrease in the natural fuel 
values of the material, and as a result, the roasting is neither so 
efficient nor so cheaply operated, owing to the need of external fuel 
for giving the required roasting temperatures. On the other hand, 
it appears to be just this class of material which is best suited for 

It is found in actual working practice that material which does 

not contain a certain proportion of silica does not work well in the 

blast-roasting or sintering processes, the resulting product being found 

to be more irregular in composition and more difficult to operate in 

the sintering plant. It would therefore appear that a certain class 

^•f fine concentrate higher in silica and lower in iron and sulphur 

intents, which is not quite so suitable for ordinary roasting (owing 

' the necessity for external heating, due to lower fuel values) is 

ininently suited for blast roasting or sintering processes, yielding 

I 'imp products very suitable for subsequent blast-furnace treatment. 

The rev^erberatory furnace thus deals most successfully with fine 
I hie concentrates high in iron and sulphur, moderately low in silica ; 
loasted, with its required flux, to the necessary extent, and then 
charged whilst still red hot into the furnaces. To relieve the rever- 
beratories of the greater bulk of the blast-furnace flue-dust, which it 



treats with more difficulty, fine concentrates, as such, require to be 
kept out of the blast-furnace charge, either by subjecting the more 
siliceous material to a preparatory sintering process, or by reserving 
the highly pyritic variety for roasting and subsequent reverberatory 


Peters, E. D., " Principles of Copper Smelting." 

Offerhaus, C, " Modern Reverberatory Smelting of Copper Ores." Eng. and Min. Journ 

June 13, 1908, pp. 1189-1193 ; June 20, 1908, pp. 1234-123G. 
Ricketts, L. D., " Experiments in Reverberatory Practice at Cananea, Mexico," and 

discussion, Trans. Inst. Miii. and Met., vol. xix., 1909-10, pp. 147-185. 
Ricketts, L. D., "Developments of Cananea Practice." Engineering and 3Iini7ig Journal, 

Oct. 7th, 1911, p. 693. 


Blast-Furnace Practice. 

Functions of the Furnacc-As Meltingr Agent— Reduction 
Smeltingr— Oxidation in the Furnace —The Pyritic 
Principle — Features of IVIodern Practice: Water - 
Jacketingr, Increase in Furnace Size, External Set- 
tling:— Constructional Details of the Furnace. 

The Functions of the Blast Furnace* — The functions of the blast 
furnace may be considered from three points of view : — 

1. As a Melting Agent. 

2. As a Reducing Medium. 

3. As an Oxidising Medium. 

In modern copper smelting practice, the blast furnace is under 
ordinary circumstances never employed in the capacity of a reducing 
medium, but is used for a variety of work in which its operations 
' ange from those of a melting furnace to those more particularly of 
an oxidising medium, as its oxidising functions are becoming 
developed to a gradually increasing extent. 

In the older processes of copper smelting, when working on 
oxidised charges, the melting and reducing functions of the furnace 
were exercised simultaneously ; when, at a later stage, sulphides were 
-melted in the charge, the directly reducing function was utilised to 
;i very much smaller extent. In the reducing atmosphere then 
maintained inside the furnace, the sulphides liquated and melted 
lown without causing much concentration of the copper in the 
product, elimination of sulphur being effected mainly by the direct 
action of heat on the pyritic constituents of the charge, and by the 
attractions between the sulphides and the oxidised compounds of 
copper present. 

When, however, increasing quantities of sulphide ore became 
available, modifications in blast-furnace smelting practice were intro- 
duced with a view to increasing the concentration of the copper, this 
J>eing attempted either by preparatory roasting or by the addition of 




oxidised cupriferous materials to the charge, sulphur being thus 
eliminated and some concentration resulting in consequence. In 
such work the furnace chiefly exercised its melting function, allowing, 
as in the case of reverberatory working, of the formation and thorough 
fusion of sulphide matte and silicate slag from the mixture of oxides 
and sulphides in the charge. In the latest developments of practice, 
the oxidation has been carried out to a continually increasing extent 
by the air blast at the tuyeres of the furnace. 

1. The Melting Functions of the Blast Furnace. — The 
blast furnace is under ordinary circumstances, usually regarded as 
the cheapest of melting agents. Compared with the reverberatory, 
the heat in the blast furnace is utilised more efficiently. Rever- 
beratory working involves the passing of a flame over the surface of 
the charge, and the transference of this heat through the mass depends 
upon the conducting power of the material itself, which is, however, 
usually poor. Although the modern reverberatory practice of melting 
thin layers of pre-heated charge both from above and from below has 
greatly increased the efficiency of the furnace in this respect, the 
closer contact of charge and fuel in the blast furnace allows of a more 
thorough communication of the heat. 

The principal features of blast-furnace working which tend to 
make it the cheaper and more efficient agent for the treatment of 
cupriferous materials — with the exception of fines — are those of con- 
struction, working, and fuel economy. 

(a) The construction of the furnace is comparatively simple, and 
it is not excessively expensive to erect ; furnaces and accessory plant 
can be purchased complete and easily set up and taken down again 
when required. 

{b) The furnace is elastic in its operation, especially where the 
supply of material varies from time to time, involving changes in the 
composition of the charge. 

(c) The furnace is readily started, shut down, and restarted at 
will, and without much difficulty or additional expense. 

(d) The operation and smelting are rapid and cheap, the capacity 
can be made enormously large ; all classes of material — except fines — 
such as ores, slags, and residues, which accumulate to a considerable 
extent round a smelter, can be conveniently dealt with directly, whilst 
fines can now, where necessary, often be prepared into a suitable form 
for blast-furnace treatment. 

(e) The heat is more efficiently communicated to the individual 
parts of the charge, in consequence of the more intimate contact of 
charge and fuel. 



(/) The fuel consumption is low, the natural fuel values of the 
iron and sulphur on the charge can be utilised, and the degree of 
oxidation (and consequent concentration) can be controlled in the 
furnace operation. 

(g) The furnace works continuously (in modern practice the 
reverberatory furnace is also continuous in its action). 

Owing to the great elasticity in blast-furnace operation, and its 
capability of dealing with practically every class of copper-bearing 
material in lump form, modern practice is of the most diverse 

2. The Blast Furnace as a Reducing Medium* — In modern 
smelting practice, with but a few exceptional instances, a distinctly 
reducing atmosphere is avoided as far as possible. This arises largely 
from the fact that the material available in modern work usually 
demands oxidation in order that satisfactory concentration may be 

In the early days of copper smelting, however, the reducing action 
was the chief function which was exercised, mainly because at that 
time oxidised ores constituted an important part of the charge, and a 
reducing action was required to obtain marketable products from such 
material. At a later stage in the development of blast-furnace practice, 
the sulphide ores which became available were roasted, and the resulting 
oxidised products were subjected to reduction smelting, in order to 
extract the metal. On such oxidised charges, blast furnaces were 
almost universally employed, using carbonaceous fuel either in the 
form of coke or charcoal, this material fulfilling the double purpose 
of fuel and reducing agent, the excess carbon causing the reduction 
of the metal from the oxidised ore. 

This operation was known commonly as " black-copper smelting. '* 
At the present time such oxidised ores are rarely met with in sufficient 
juantity by themselves to be worked by this method, which involves 
ilso very serious losses in operation. Further, such oxidised materials 
are in many cases valuable for smelting along with sulphide charges, 
greatly assisting the concentration, and it is usually advantageous to 
employ them in this manner. 

The losses and difficulties in "black copper smelting'* are, how- 
ver, of interest in so far as they apply to certain analogous problems 
in modem work. These difficulties in reduction smelting arose largely 
^rom three causes : — 

(a) Losses of copper in the slag. 

(h) Simultaneous reduction of iron with the copper, 

(c) ChiUing in the furnace hearth. 


{a) In the case of reduction smelting where sulphides are not 
present in any appreciable quantity, the losses of copper may be 

(i.) As silicate, or 
(ii.) As metal. 

(i) Sulphur is the natural protector of the copper in the furnace 
charge, as, owing to their powerful affinity, a fusible, fluid and dense 
product is formed, which is very slightly soluble in slag ; and on this 
account, a ready separation of the copper from the earthy materials 
can be effected. So long as sulphur is present in moderate quantity 
there is little chance of copper entering the slag as silicate. 

In reduction smelting, however, and especially in black copper 
smelting where sulphur is lacking, such losses are liable to occur, 
since copper oxide is itself strongly basic, and readily fluxes off with 
silica at high temperatures, yielding silicates. These products are less 
dense, and are markedly soluble in the other silicates which constitute 
the slag ; moreover, the copper oxides themselves are likewise partly 
soluble in, and are readily carried in suspension by, the silicate slags. 

In order to prevent such losses as much as possible, the reducing 
conditions in the furnace must be increased by the employment of 
more coke, so as to ensure the reduction of the copper oxides and 
silicates. These reducing conditions must not, however, be too drastic, 
especially if the temperature of working be high, on account of the 
great tendency to cause (b) a reduction of metallic iron, which results 
in the formation of bears and scaffolds, with their attendant difficulties 
of removal and their interference with working. 

Between these opposing causes of loss and difficulty, a careful 
balance has to be observed in the .smelting operations. (In modern 
practice, losses of copper as silicate and oxide, for reasons such as 
those detailed above, occur to a marked extent in those operations 
where the sulphur is present in small proportions only, and particularly 
where the reactions are intensely oxidising, as in the furnace-refining 
operations and the later stages in the conv^erter process. The slags in 
such cases usually carry considerable quantities of copper in the form 
of silicate and oxide, not infrequently to the extent of 20 to 30 per 
cent., or even more. The quantity of this slag is, however, kept as 
small as possible, and copper in the material is readily recovered by 
the addition of these slags to the blast-furnace charge.) 

(ii.) Losses of copper as metal also, were formerly serious in black- 
copper smelting, the metallic copper held in suspension in the slag 
being indeed the chief source of loss in this method. The efficient 



separation of copper from slag, especially in the small quantities 
formerly operated, was therefore of importance. Satisfactory settling 
was. however, difficult of apphcation, since the behaviour of metallic 
copper is very different from that of sulphides. It is much less fusible, 
much less fluid, and the small globules, as reduced, do not readily 
coalesce, whilst the high temperatures favourable to good fluidity of 
the products and to good setthng, promote copper losses from the 
other causes noted above. 

Moreover, the high melting point of the metal and its great con- 
ductivity added to the difficulties in providing suitable arrangements 
for settling, since the copper not only tended to chill readily in any 
external settler, but it was also very liable to do so in the crucible of 
the ordinary form of water-jacketed blast furnace, such masses being 
exceedingly difiicult to remove, whilst the working of the furnace was 
necessarily much interfered with. 

In order to conduct the necessary internal settling, the older type 
of blast furnace was required, in which water- jacketing near the hearth 
was dispensed with, a large crucible bottom of non-conducting brasque 
or brickwork being employed instead. Such a form of furnace is not 
adapted to the modern methods of smelting where enormous capacity 
and output are essential, whilst such a system of working interferes 
with the rapid and continuous smelting of large quantities, to a greater 
extent than if the whole of the molten products are run out of the 
furnace continuously and the settling performed in an external vessel. 

3. The Blast Furnace as an Oxidising Medium : Sulphide 
Ores in the Blast Furnace. — In modern blast-furnace practice, the 
oxidising function of the furnace is the principal feature of working. 
Sulphide ores now constitute the chief source of copper, and the 
smelting operations involve the oxidation of the accompanying con- 
stituents and the elimination of the resulting oxidised products. 

Such ores when smelted in the blast furnace with carbonaceous 
fueL and under the reducing conditions characteristic of the older 
methods of working, would yield a product showing low concentration 
of the copper, since the reducing conditions would largely retard the 
oxidation of sulphur which is an essential for the enrichment of the 
matte. Except for the sulphur eliminated from the pyritic con- 
stituents by the direct action of heat, and a certain quantity by the 
interactions with oxides as already indicated, the loss of sulphur would 
be slight. The furnace under such circumstances would thus tend 
mainly to exercise its melting function, and the result of such 
working would be the melting down and subsequent separation of 
the sulphides and slag, with even less tendency to concentration 


than occurs in the reverberatory furnace, where the atmosphere is 
less distinctly reducing. 

The modern method of smelting sulphide ores being essentially 
an oxidising process, it is necessary that oxygen be added to the charge 
with the object of promoting the elimination of the sulphur and iron, 
and the consequent concentration of the copper. 

This oxygen may be added in one of three ways : — 

A. Addition of oxygen to the charge previous to the blast furnace 
smelting operation (Roasting). 

B. Addition of oxygen to the charge during the smelting operation 

i. By adding oxidised materials to the charge (Blast-furnace 
smelting with carbonaceous fuel). 

ii. By using the air blast of the furnace for oxidising the iron 
and sulphur, thus at the same time utilising these 
elements as fuel and proportionately diminishing the 
amount of carbonaceous fuel required (The pyritic prin- 
ciple of smelling). 

A. Roasting practice has already been discussed, and the reasons 
for avoiding the operation where practicable, on account of the 
expenses of an extra process, the losses involved, the fineness of the 
product, and the loss of fuel values, have been indicated (Lecture IV., 
pp. 66-80). 

B i. Addition of Oxidised Charges in the Blast Furnace. — The 
tendency for oxidised cupriferous materials to interact with sulphides 
finds useful application in copper smelting, since it assists the 
concentration of the copper in the resulting mattes. The principal 
reactions involved in this method are — 

2CuO + Cu,S -> 4Cu -h SO. 
2CU2O + CugS -> eCu + SO2 
CUSO4 + Cu^S -> 3Cu + 2S0, 

whereby copper is produced and sulphur is eliminated as SO.,. The 
liberated copper interacts with the excess of iron sulphide usually 
present in the furnace charge, and enters the matte as sulphide, whilst 
the iron which is thus set free is oxidised and carried into the slag as 
silicate, the ultimate reactions being indicated approximately by the 
equation — 

2Cu + FeS 4- a;FeS -> CugS . xFeS (matte) + Fe (oxidised and enters slag). 



Copper silicates readily interact with iron sulphides in the charge, 
producing copper sulphides and iron silicates, thus — 

C112O . xSiOa + FeS -> CugS (enters matte) + FeO . ajSiOa (enters slag). 
6(CuO . arSiOo) + 4FeS -> 3Cu.S (enters matte) + 4(FeO . .TSiOo) + 2.rSi02 (enter slag)+S02. 

All the above reactions lead to an enrichment of the matte in 
copper contents, and at the same time, to the transference of iron 
from the matte to the slag, and although the conditions in the more 
reducing atmosphere of the coke-fed blast furnace are not so favour- 
able to the fullest operation of these reactions as are the more neutral 
conditions of the reverberatory, the addition of oxidised materials 
constitutes a valuable means of increasing the concentration in this 
method of smelting. 

The blast furnace is thus also particularly suited for the recovering 
of the copper from the oxidised residues, such as converter slags and 
scrap, " calcine-barrings,'' and the like, which accumulate in very 
considerable quantities at a smelter, and which by reason of their 
carrying much copper as oxide or silicate, not only add their quota 
of copper to the products, but materially assist the concentration and 
the furnace operation generally. 

B ii. The Pyritic Principle in Blast-Furnace Smelting. — This is 
the most important principle introduced into modern blast-furnace 
smelting practice. 

It has been evolved by the application of the results of experiments 
conducted from two different points of view — one series mainly on a 
laborator}^ scale, the other from actual industrial practice. 

Starting from theoretical considerations, John Holway demon- 
strated by experiment that the heat of oxidation of the iron and the 
sulphur of pyritic copper ores was so great as to make their smelting 
a self-supporting operation under suitable conditions. On the other 
hand, within comparatively recent years, smeltermen as a result of 
working practice, have found that an increase of sulphides on the 
furnace charge has led to less and less carbonaceous fuel being 
necessary for the smelting operations, providing that the conditions 
in the blast furnace be sufficiently oxidising. 

In utilising these results for general blast-furnace practice, the 
extended and successful application of this pyritic principle has led 
to marked advance in modern working. 

The results obtained in a series of trials at the Keswick smelter, 
California, are typical of such experiments on a practical scale, and in 
spite of the two anomalous instances, the general effects of the increase 
of sulphides in the chai^ are strongly marked (see Table IX., p. 120). 



TABLE IX. — Effect on Coke Consumption of Increased Sulphur 
IN the Furnace Charge (Keswick Smelter, Cal.). 

Sulphur in Charge. 

Coke Consumption. 

6-8 per cent. 

15-7 per cent. 

7-7 „ 

16-3 „ 

13-6 „ 

10-2 „ 

17-0 „ 


19-5 „ 

8-5 ., 

22-8 „ 

7-1 „ 

24-5 „ 

C-8 „ 

Recent practice at Anaconda affords another instance of the 
utilisation of the pyritic principle. A large quantity of the ore 
available (known as second-class ore) requires wet dressing before 
it can be treated most profitably at the furnaces, and the operation 
thus produces considerable quantities of sulphide concentrate, of 
which a moderate proportion is coarse — well suited for blast-furnace 
treatment. The charge if submitted to reduction smelting with 
carbonaceous fuel, would yield a matte too low in copper contents 
for immediate converter treatment, since there is not available a 
sufficient supply of oxidised cupriferous material to effect a high 
enough concentration for the direct production of a converter-grade 
matte. Instead of roasting so as to reduce the sulphur contents to 
the required degree, and then smelting with the usual amount of 
carbonaceous fuel, the pyritic principle has been utilised to the 
fullest possible extent, by smelting the raw charge containing as 
much of the coarse concentrate as is available, with a strongly 
oxidising blast, thus effecting the desired concentration, and occa- 
sioning the use of a lower coke proportion than would otherwise 
have been necessary. By gradually increasing the sulphide on the 
charge until the sulphur proportion reached 8 to 9 per cent., the 
coke consumption was reduced to about 11 to 12 per cent. During 
the past two or three years the advantages of introducing more and 
more sulphide have become so apparent, that increasing quantities 
of I inch concentrates are being included in the charge, and 
although such material is exceedingly difficult to deal with in 
the blast furnace, the advantages arising from its use outweighs 
the trouble it causes in actual working. By this further increase of 
the sulphur proportion, from the former 8 to 9 per cent, up to II to 12 


per cent., the coke consumption has been steadily reduced until it 
now amounts to about 9 per cent. onty. 

The fuel value of the iron and sulphur is augmented at a rate 
much greater than their actual increase in numerical proportion would 
suggest, on account of the much higher calorific intensity of large and 
massive quantities of fuel burned at once than that resulting from 
smaller amounts disseminated throughout a mass of inert material 
such as gangue. 

The practical application of the pyritic principle to blast-furnace 
practice thus involves the employment of the furnace as a medium 
for conducting the required oxidation of the charge, as a result of 
which, the heat of this combustion proportionately reduces the amount 
of carbonaceous fuel required for the smelting and separation of the 
products, whilst at the same time the desired concentration is also 
effected. The basis of such working is, therefore, the powerful oxidising 
action within the furnace itself, and the fullest utilisation of the heat 
resulting from this oxidation of the sulphides. 

In order to supply the heat necessary for the reactions and fusions 
of smelting, a definite quantity of fuel is essential in the furnace. In 
those cases where the proportions of sulphide are not sufficient to 
supply the required amount, a supplementary quantity of coke fuel 
becomes requisite. 

The extent to which coke is necessary for the smelting operations 
decides whether the process may be termed " true pyritic '' or " partial 
pyritic " smelting. In the former case, the coke allowance may be 
reduced to such small proportions that its influence in the smelting 
zone of the furnace is practically negligible. 

In partial pyritic smelting, coke is necessary to the extent of sup- 
plementing the heat derived from the sulphide fuel, and the proportion 
'•mployed in modern work is reduced to the lowest possible quantity. 
Xot only is economy in coke allowance one of the chief essentials 
in furnace management, but the presence of a larger amount than is 
absolutely necessary decreases the efficiency of the smelting operations, 
nince, owing to its reducing action and its consumption of the oxygen 
in the air blast which is to be utilised for the combustion of the iron 
and sulphur, the concentration of the copper in the resulting matte 
Aould be decreased. 

The extent to which the pyritic principle may be operated in 
actual working depends in the first instance upon the nature of the 
charge itself, especially upon the relative proportions of copper, iron, 
and sulphur, and on the quantity of gangue. Since these vary 
in the ore supply of different localities, the extent to which the 


principle may be applied and the coke consumption be reduced, 
will be subject to alteration accordingly. 

Thus in the case of an ore which contains such proportions of 
these constituents as would on simple melting yield a matte of con- 
verter grade, the pyritic effect in the furnace would necessarily be 
very small, and the smelting would be almost entirely a melting 
operation requiring from 10 to 15 per cent, of coke on the charge, 
even though the sulphur contents of the charge be high. Ores and 
charges of such a composition are, however, rarely met with in modern 
practice, the ratio of copper to iron sulphides usually being low. 

On the other hand, in the case of an ore consisting largely of iron 
sulphides with but little copper — i.e., a massive low-grade pyritic 
ore — the pyritic effect in the furnace might reach a maximum, and 
the coke required on the charge be reducible to very small propor- 
tions. Such material is well-suited for true pyritic smelting. 

Hence modern practice ranges from the true pyritic smelting, 
where pyritic fuel is principally employed, through varying degrees 
of partial pyritic smelting, where the pyritic fuel is supplemented to 
the required degree by coke, to reduction smelting, relying mainly on 
carbonaceous fuel for the necessary heat supply. 

In all cases, the object of the operation is to oxidise inside the 
furnace so much sulphur and iron as is necessary to yield a matte 
product of converter grade, utilising the natural sulphide fuel values 
of the material so as to reduce to the lowest possible proportion the 
quantity of coke required. 

Features of Modern Practice, — Apart from the applications of 
pyritic smelting, which will be considered separately, three features 
of great importance have been introduced into modern blast-furnace 
working. These involve : — 

A. The practice of water- jacketing the furnace. 

B. The development in the size of the furnace. 

C. The practice of external settling. 

A. The Practice of Water-jacketing. — The evolution of the blast 
furnace from the primitive hole-in- the-ground form to the modern 
type may be rapidly sketched. In its early stages, the development 
was carried out mainly on the Continent of Europe, following the 
course of the enclosing of the charge in shafts which became of gradu- 
ally increasing height, the introduction of blast through tuyeres 
near the bottom of the shaft, and the arrangements for collecting the 
molten materials in the hearth, and for tapping. By the year 1850 
a typical form of furnace was represented by the Mansfeld pattern, 

[To face page 122. 

Fig. 32,— Modern Blast- Furnace Shell of Sectioned Jackets (P. & M. M. Co.). 


which consisted of a rectangular firebrick shaft enclosed by massive 
stonework. At the lower extremity was a hearth constructed of 
refractory material, usually of brasque — a mixture of fireclay and 
coke — well tamped down. The dimensions were from about 2 feet to 
2 feet 6 inches broad, 14 feet to 16 feet high, with two tuyeres of H 
to 2 inches diameter, supplying blast at 4 to 10 inches water pressure ; 
the capacity of such a furnace being about 4 tons per twenty-four 
hours. It is of interest to note that this form of furnace possessed 
arrangements both for internal or external settling of the products, 
the usual practice being, however, to allow the smelted material to 
collect and settle in the hearth. In endeavouring to increase the 
capacity of the furnace and the rapidity of working, as well as to ensure 
efficient settling of the products, it became necessary to maintain a 
high temperature in the lower parts ; but in consequence of the exces- 
sive heat and the corrosive nature of the molten materials, the most 
refractory brasquing available was rapidly attacked, and the necessity 
for adopting means to prevent the destruction of the furnace linings 
became apparent. 

The use of water- jacketing for this purpose had long before been 

pplied to certain branches of cast-iron refining, and in 1875 the 
i^iltz water-jacketed blast furnace was introduced for the smelting 
of lead ores. This form of furnace was circular in horizontal section, 
and the boshes consisted of two concentric shells between which a 
stream of water circulated. This principle was quickly adopted for 
the purposes of copper smelting furnaces, although modifications were 
found to be necessary in certain particulars before perfectly successful 
working was achieved. Owing to the higher temperatures prevailing 
in the furnace, the height to which the water-jackets were carried 
required to be increased, and it was chiefly when the rectangular form 
of furnace was introduced that the thoroughly successful application 

f water- jacketing was accomplished. This feature in blast-furnace 
work was rapidly and very successfully developed by the American 
copper smelters when the new estabfishments in the West were opened 
"p. and the substitution of the older form of fining by metallic water- 

(X)\ed jackets, which in comparison are practically indestructible, 
immediately led to an enormous improvement in smelting practice. 

The modem blast furnace is essentially a water- jacketed shell from 
charging floor to base plate, rectangular in plan, and completely 

Many of the advantages of such a furnace construction are apparent, 

'nd have been referred to in discussing the furnace as a melting agent. 

I'hf salient features of the modem water- jacketed fumace are : — 



(i.) Water- jacketed furnaces are planned, constructed, and erected 
simply and with ease. 

(ii.) The first cost of the furnace, making allowance for excavation 
and foundations, is not unfavourable to the Avater- jacketed furnace, 
whilst the ease of fitting and the interchangeability of parts due to 
sectioning, reduce the costs of erection. 

(iii.) The convenience and simplicity in operation of the water- 
jacketed furnace are very marked, whilst the permanence in the shape 
tends to greater uniformity of working and to ease of management. 

(iv.) Accretions and the general difficulties of working are readily 
dealt with and controlled, barring and other operations being more 
conveniently conducted. 

(v.) The repairing of water-jacketed furnaces is rendered very 
simple, cheap, and rapid in operation, the principle of sectionising 
allowing of the ready removal or replacement of the jackets for repairs ; 
the saving in time, labour, and general expense being particularly 

(vi.) The elasticity of the furnace, both as regards size and 
management, has been enormously increased, and the successful 
extension and working of the large modern furnaces have only become 
possible with the adoption of this feature. 

(vii.) Water-jacketing has allowed of the rapid driving of furnaces, 
leading to an enormous increase in the output per square foot of 
hearth area, by permitting intense heating inside the furnace, and 
rapid withdrawal of the molten products. 

The chief consideration affecting the adoption of water-jacketing 
in any locality might be the scarcit}'^ or unsuitability of the water 
supply, which may necessitate a choice between the employment of 
brick furnaces, or the crushing, roasting and reverberatory treatment 
of the ore. In cases where the water supply is not well suited for 
jacketing purposes, settling or other preliminary treatment of the 
water might be required. 

The former objection to water- jacketing on the assumption of 
valuable heat being carried away by the jacket water, thus involving 
a waste of fuel, has proved to be groundless in practice ; with good 
management such heat losses are smaller in amount and less damaging 
in effect than those due to radiation from highly heated brick walls, 
quite apart from the actual necessity for such jacketing in modern 
furnace construction, even had such losses been marked. 

B. The Development in Furnace Size. — The blast furnace increased 
but slowly in size during the nineteenth century up to 1850, and the 
dimensions of the most advanced type did not exceed 4 feet by about 


2 feet 6 inches internally at the tuyere level, the capacity being about 
4 tons per day. Furnaces at this period were usually square or circular 
in section. 

The size of such furnaces was largely dependent on the penetrating 
power of the blast, and a slight increase in cross-section resulted 
gradually, as improvements in the mechanical contrivances for pro- 
ducing blast were developed. This, however, soon reached a limit, 
owing to the difficulties in making the blast penetrate to the centre 
of the charge in the wider furnaces, and to the disproportionate cost- 
liness and increased working difficulties attendant on such practice. 
It was further found that the high pressure required in order to force 
the blast through an increased width of charge produced an intense 
local heating effect against the tuyeres, resulting in high slaor losses 
and low concentration on smelting, whilst the consumption of fuel was 
much increased. 

An important modification in blast-furnace design was introduced 
in 1863, when the principle of increasing the size of the furnace in 
direction of its length, whilst maintaining the width which had been 
found best suited to economical working, was applied by Raohette. 
This was first intended for the purposes of lead smelting, but the 
principle was quickly recognised as having important applications 
to copper smelting practice, and was readily adopted and developed. 
It has become the basis of all subsequent modern copper blast-furnace 
design, and the gradual increase in dimensions up to the enormous 
blast furnaces with huge outputs of the present day has been made by 
xtending the length whilst maintaining a relatively small width. 

For some time development proceeded along these lines slowly 
and with much caution, chiefl}^ owing to the difficulties anticipated 
in the management of such large units. Up to 1885, the largest blast 
furnace (at the Parrott Smelter, Butte) was but 8 feet long by 
36 inches wide; by the year 1900 the dimensions had reached 10 feet 
by 42 inches. Subsequently, under the direction of the remarkably 
enterprising management of the Washoe Smelter at Anaconda, a 
^wonderful era of furnace extensions was commenced, and is indeed, 
ill undergoing development. 

Here in 1902, blast furnaces 15 feet long by 56 inches wide were 

;ted, the plant eventually consisting of seven such furnaces built 

straight line, and situated 21 feet apart from each other. A 

>ly augmented ore supply subsequently coming to the smelter 

'treatment, an increased furnace capacity was required, for which 

a very limited suitable space was available. Mr. E. P. Mathew- 

m, the smelter superintendent, determined upon attempting the 




revolutionary idea of joining up two of the 15-foot furnaces by 
bridging over the 21-foot space between them, and continuing the 

vertical side water-jackets across this space, thus forming a furnace 
15 -j-21 -fl5, or 51 feet in length. No work on such a large and 


boldly conceived scale had ever been attempted before, and many 
difficulties in construction and operation were anticipated. 

Mathewson first conducted a series of constructional trials, and 

found in the first instance that by taking suitable precautions, it 

would be possible to carry out these changes whilst the furnaces 

themselves were running. It was found that it was possible to 

remove or replace single jackets without shutting down the furnace, 

j by the device of forming a crust against such a jacket, of sufficient 

thickness to bear the weight of the charge for the short period of 

time during which the change was being made. Such a crust is 

j readily obtained by shutting off the tuyeres in the particular jacket 

' and in its neighbours, and maintaining a rapid stream of cold water 

through these jackets. Further, it was found that any desired portion 

\\ of the sides or hearth of such a long furnace could be well barred 

' and cleaned whilst the rest of the furnace was in operation, whereas 

such barring and cleaning on a small furnace seriously interrupted 

j; the working, and reduced the capacity. 

The prehminary tests being satisfactory, the necessary construc- 
I' tional work was carried out whilst the two furnaces were in blast ; 
\ the inner end jackets of these furnaces were taken down, and in a 
short time the new 51-foot furnace was in regular operation, and 
proved so remarkably successful that two other pairs of furnaces 
jl were similarly joined up. In the following year a still further great 
■ extension was made by joining up in a like manner the end 51 -foot 
< furnace to the last remaining 15-foot furnace, by again bridging over 
the intervening 21-foot space, thus constructing a furnace of the 
enormous length of 51 + 21 + 15, or 87 feet. 

It was at one time intended to carry this progress still further by 
i| joining up the other two 51 -foot furnaces, so as to make a single one 
'23 feet in length, but certain difficulties in the matter of bringing 
>ke supplies to the two sides, under the special conditions of available 
i floor space, and the disastrous effects of the financial panic of October, 
1907, stopped all extension work for the time. Such extensions would 
however, present no real difficulties either in construction or in sub- 
sequent furnace management or operation. 

Figs. 35 and 36 indicate in plan and elevation the arrangement of 

^he plant and accessories for these extended furnaces. Each 15-foot 

irnace had its own settler situated in front, and these have been 

ifetained without any change of position or any further additions. 

The liearth of the newly bridged portion slopes from the middle of the 

ridge, to the tap-holes of the old furnaces, which still serve this pur- 

•f)*'f for the larger ones, and from which a continuous stream of matte 



and slag flows through a slag spout to the settler in front. The sidef 
water-jackets of the old furnaces remain, being built up in two sets 

Fig. 35. — Plan of 51-foot Blast Furnace, Anaconda, indicating Position of Crucibles, 
Spouts, and Connecting Bridge between Old Furnaces. 

Fig. 36.— Longitudinal Section and Part Elevation of 87-foot Blast Furnace, Anaconda, 
indicating Crucibles of Old Furnaces^ Bridge, and Jacketing. 



of panels, each 7 feet 6 inches wide, whilst the new bridge portions 
are constructed of three sets of jackets, each 7 feet wide. 

The furnaces in their lengthened form have proved a tremendous 
success, far indeed beyond the anticipation of the designers and 
managers. This is largely due 

(a) To the increased efficiency and economy of replacing a number 
of smaller furnaces situated end to end by a single large furnace ; 

(b) To the increased intensity of heat and reactions owing to large 
massed quantities of fuel burned at once, and to large masses of 
material being smelted and in a state of chemical activity. 

The advantages which result from such lengthening of blast fur- 
naces are : — 

(i.) Gain in hearth area without extension of the blast-furnace 
floor and building. 

(ii.) Increase in smelting or hearth area and in consequent capacity, 
at a rate very much superior to the extra water- jacketing involved. 
Thus, in the 51 -foot furnace, the capacity has been increased in the 
proportion of 3-8 to 1, the jacketed surface has increased only at 
the rate of 2-4 to 1. The output has increased at a much greater speed 
than was actually anticipated from the additional hearth area. 

(iii.) A very marked saving of fuel. The amount of coke required 

for similar charges has been reduced by one-tenth ; more than 1 1 per 

ent. was required formerly on a charge, only 10 per cent, was neces- 

iry under the new conditions. 

(iv.) The rapidity of working of the furnace has increased owing 

to the effect of the narroAv width and small crucible dimensions as 

compared with the length. This has caused a more rapid flow through 

the furnace slag-holes, thus preventing the formation of obstructions, 

and tending to wash out any which might threaten to stick. 

(v.) Higher furnace temperatures result, and both slag and matte 
re hotter than in smaller furnaces. In consequence more siliceous 
lags can be run. thus saving the cost of the fluxes which might 
therwise be necessary. 

(vi.) Marked decrease in incrustation. Crusting is most hkely to 
ccur at points where the smelting activity is lowest, and in the cooler 
parts of the furnaces, such conditions being usually prevalent at the 
^^oniers, where the shape also assists in the holding up of material, 
^^^usting is one of the chief troubles to be prevented and overcome 
^^H operating the blast furnace. 

I^r The elongated furnace of 87 feet length practically takes the place 
f five shorter ones, representing no less than 20 comers and 10 end 
ackets ; the new furnace thus reduces the opportunities for crusting 




at least five-fold. In this way the hearth area has been very greatly] 
increased, with still but two ends to hold crusts. The long furnace- 
walls with their ends so far apart, in addition, offer much less oppor- 
tunity for the formation of crusts than do the side walls of shorter 
furnaces, accretions obtain little support, and often tend to break 
down under their own weight, whilst they can be more readily' 
removed by barring, on lowering the height of the furnace charge for 
a time. 

(^ai.) The elasticity of the furnace operations has been much' 
increased. In short furnaces, cleaning and barring for the removal,, 
of obstructions, etc., necessitate the shutting down of the unit, oftenj 
a complete taking down of the furnace-walls and their subsequent! 
replacement, followed by a re-starting of the furnace work. The ideal; 
in modern work is continuous running of the unit. The larger furnaces; 
allow of such practice, since they can be kept in operation whilst a 
particular portion is undergoing cleaning or repair. As stated above, 
the elongation of the furnaces themselves was conducted whilst the 
older 15-foot portions were working. Leaky or worn-out jackets or 
spouts are readily removed without serious interference with the 
working of the rest of the furnace, and this operation usually 
requires a few hours only. 

(viii.) The charge may be varied in different parts of the furnace 
to suit special requirements, without interfering with the general 
operations. Thus, suitable additions for the smelting out of crusts, 
or variations in the charge to reduce corrosion near the 21-foot bridge, 
can be effected whilst the furnace is running as usual. 

(ix.) Increased flow of material through the settlers is effected 
without decreasing the efficiency of the settling. Each settler now 
serves 25 feet of furnace-hearth length, instead of the 15 feet of the 
smaller furnaces, and in spite of the more rapid passage of the 
materials, the settling is actually better and the resulting slag cleaner, 
owing to the higher temperatures of working and the consequent 
greater liquidity of the products, whilst the settler is also hotter. 
Thus the greater output of material has required no extra labour or 
construction on the tapping floor, though tappings are now more 

(x.) The labour costs per ton of furnace capacity are greatly i 
reduced, as are also the operating and management costs, since such: 
labour and control are to a large extent dependent on the number of 
units comprising the plant. 

(xi.) The initial cost, per ton of furnace capacity, is also much; 
reduced. In the elongated furnace, the settlers have not been added; 


to, the old slag notches only are required to do duty as before, and the 
older equipment for bracing and trussing provides for much of that 
required in the extensions whilst the original building itself served 
for the housing of the increased furnace area. 

(xii.) Further extension of the furnace length is readily possible 
if desired. 

The older 15-feet furnaces had a smelting capacity of 5-6 tons 
per square foot of hearth area per day, those of 51 feet length smelt on 
an average 6-72 tons per square foot daily, whilst the output of the 
87-foot furnace amounts to 3,000 tons of material daily, corresponding 
to 3,000 -f- 87 feet x 4 feet 8 inches, or about 7-5 tons per square 
foot of hearth area. Whilst this particular smelter is of course unique 
in the dimensions, equipment, organisation and management of its 
plant and the magnitude of its operations, and though at most 
modern smelters the ore supphes and smelting conditions do not 
.admit of the introduction of such enormous units ; at the same time 
the principles which underhe the great advantages of the longer form of 
blast furnace have had an important influence on blast-furnace 
equipment and design generally. The constructional details of these 
large furnaces are, for the most part, common to all modern blast 
furnaces ; it is mainly the size and capacity which are exceptional. 
The usual length adopted at smelters with more modest output varies 
from about 15 to 25 feet, with a smelting capacity of from about 400 
to 800 tons per twenty-four hours, depending naturally on the working 

C. The Practice of External Settling. — In connection with modern 
l)last-fumace practice, the feature of external settling is of much 
importance, its adoption having had a marked influence on : — 

{a) The efficiency of separation of the smelted products, and 

the production of clean slags. 
(h) The output, and rapidity of working of the furnace, 
(c) The control and organisation of the smelting processes. 

(a) The function of the blast-furnace plant is the concentration 
f the values into a matte of correct grade for further treatment, and 
If- production of a slag which is sufficiently clean — that is, free from 

Ipper and other values — to allow of its being disposed of as waste, 
imediately. Numerous factors decide the copper contents of the 
ig which is economically the cleanest — the general average is about 
25 to 0-35 per cent, of copper. The actual condition of the copper 
the slags is a matter of some uncertainty, and it does not appear 
■mprobable that very small quantities of sulphides may actually be in 


solution in the silicate slags. The general concensus of opinion, how- 
ever, favours the view that much of the copper which is present exists 
in the form of minute shots of the matte, actually held in mechanical 
suspension, and this is certainly the case when the copper contents 
exceed the limits stated above. In consequence, it is frequently noted 
in practice that the copper in the slag increases with the grade of 
the matte. The question has been reviewed by L. T. Wright who 
suggests some actual solubihty of matte-products in the slag. 
Wright's curve indicating the connection between matte-grade and 
slag values is reproduced in Fig. 37. This connection might how- 
ever, possibly result from the fact that the individual shots of matte 
are themselves higher in copper contents, since it may be assumed 
that in fairly clean slags practically the same number of shots are 





20 25 30 35 40 

Percentage of Copper in Matte. 
-Copper Contents in the Slags accompanying Mattes of Various Grade. 

held up, owing to the forces of capillary attraction and surface tension, 
and that the increased density of the higher grade mattes would 
influence but slightly their downward settling when in such a fine state 
of division.* 

The molten products of the blast-furnace operation are separated by 
the settling of the matte and slag under the action of gravity, and the 

* Later work on this subject has been published by W. Wanjuko£f : — " Investigations on 
the Conditions governing the Entry of Copper into the Slags on Matte-Smelting, on the 
Chemical Form in which such Copper exists, and on the Lessening of the Copper Losses in 
Slags." Metallurgie, 1912, Vol. x., Nos. 1 and 2, pp. 1-27, Jan. 8 and 22, 1912. 



production of the economically cleanest slag depends upon the fulfil- 
ment of those conditions which allow of the most perfect dowaiward 
setthng of the small particles of matte. The three main requirements 
for efficient settling, apart from the composition of the slag, are : — 

(i.) Sufficiently high temperature. 

(ii.) Opportunities as regards time, rest, and space for quiet 

(iii.) Large masses of heated products. 

In each of these essentials, the method of external setthng, as now 
conducted at modern smelters, best satisfies the conditions required 
for successful work. 

The present practice is to make no attempt to conduct settling in 
the blast furnace, but to run the products through and out of the 
furnace with the greatest speed attainable, aiid to allow the matte 
and slag sufficient time and opportunity to settle and separate in some 
independent and external vessel, which stores the matte and allows 
the clean slag to run straight away to waste. 

The former method of inside settling gave rise to many difficulties 
in practice, but objections were urged against the external settler, to 
the effect that heat might be wasted by the abstraction of hot 
materials from the furnace to an exterior vessel, and that the settling 
would not be efficiently conducted outside, as in the very hot interior 
of the smelting furnace. Modern practice has proved conclusively that 
both objections are groundless. Such heat as is carried away by the 
continual stream of molten material can usually be well spared in the 
modern plant, which is driven so rapidly that an abundant supply of 
exceedingly hot matte and slag pass through to the settler, whilst the 
results of every-day working demonstrate the efficiency of the external 
settler, which cannot be equalled, far less surpassed, by any method of 
inside settling, under modern smelting conditions. Thousands of tons 
of slag pass daily through the settlers, clean enough to discharge 
straight to the dump, the copper contents rarely exceeding 0-40 per 

(b) The modern conditions of rapid working and large output 
render the use of external settlers practically essential, owing to the 
flouble work of smelting and separating being no longer confined 
o one and the same vessel. The aim in present practice is to exercise 
e smelting function only of the furnace, and to do so to its fullest 
pa<jity, smelting for matte of the desired grade as rapidly as possible, 
and therefore running the products through the furnace in a con- 
stant rapid stream and allowing them to settle quietly outside. Under 




these circumstances the furnace itself smelts most economically and 

It will be recalled that present-day practice involves the sub- 
sequent treatment of the fluid matte — product in the converter, so 
that whilst the former methods of working might have possessed certain 
advantages for the settling and storing of matte in the small furnaces, 
and then tapping out and casting into cakes for subsequent treat- 
ment, such methods have practically no application to modern 
systems of working. 

Internal settling almost invariably leads to the accumulation of 
debris, of chills and of any infusible masses of material which may be 
produced in the furnace, occasioning delay in the operations, waste 
and difficulty in working, and so interfering seriously with the speed 
and continuity of the smelting, as well as decreasing the output of: 
the furnace. On the other hand, a rapid flow of hot molten material 
through the furnace not only tends to prevent this formation of chills 
or accretions, but greatly assists in the dissolution or removal of such 
as might be formed. Should the production or collection of such 
masses be transferred to the settler instead, they are more readily 
attacked and remedied without interfering with the continued 
operation of the furnace. 

Further, the nature of the hearth which would be most satis- 
factory for internal settling is not at all suited for modern smelting 
conditions. The ordinary water- jacketing would have too marked a 
cooling effect on the hearth for the materials to remain sufficiently hot 
and fluid to allow of proper settling, whilst a brasque or similarly 
lined hearth suitable for such settling would, under the present 
conditions of rapid driving and intense reactions, be unable to 
withstand the highly corrosive and abrasive action to which it would 
be subject, so that breakouts, necessitating delays and repairs, would 
constantly occur. Water- jacketing in this portion of the furnace is 
indeed an essential for modern conditions, and consequently rapid 
driving and quiet internal settling in the same area are quite incom- 
patible. The modern fore-hearth, on the other hand, is accessible 
and easy of repair, and in the event of any trouble occurring 
therein, the furnace itself can continue its smelting activity to the full, 
since other suitable arrangements can readily be made for tem- 
porarily dealing with the products. 

(c) The functions of the blast furnace in the modern smelting 
scheme are particularly dependent upon the employment of the external 
settler in conjunction with it. The work of the furnace plant is to 
produce as rapidly as possible, a supply of suitable grade matte for the 

.5 03 
O bO 

.s I 

^-' 3 

0) O/ 


i s 



'"5 . 




converters ; large quantities of hot fluid matte must be available at 
a moment's notice, and such demands are often very erratic, being 
dependent on the working of the converter plant and the refining 
furnaces. It is essential to the successful operation of the blast furnaces 
that the manager should be in a position to work his furnace as rapidly 
and continuously as possible, which is best attained by making 
the output independent of irregular tappings of matte just when 
required by the converter department. The settlers, in exercising the 
function of reservoirs for matte, from which the converter department 
may draw at will, allow of regularity of working and rapidity of out- 
put in a manner possible in no other way. The only alternative, using 
internal settling, would consist of tapping out matte at regular intervals 
and casting such material when it is not immediately required, a 
wasteful and unnecessary practice incompatible with modern ideas of 
smelting work. 

During the early stages of the development of smelter plant, the 
use of reverberatory fore-hearths received considerable attention, the 
principle being to build a fire-box in communication with the settler, 
so as to ensure a sufficient supply of heat in the vessel for efficient 
setthng. Modern furnaces however, usually supply a large enough 
quantity of very hot and fluid matte and slag as to allow of very 
efficient separation without the use of extra heating, providing the 
position and construction of the settler is suitably planned, as will be 
described in due course. 

The Construction of the Blast Furnace. 

Dimensions. — The modern blast furnace is a long, narrow, water- 
cooled shell, rectangular in plan. The dimensions, particularly the 
length, vary greatly, being regulated according to the anticipated 
output of the furnace -unit. The size is generally expressed in terms 
of the internal dimensions at the tuyere level, which represents the 
smelting area. The width of the modern furnace varies usually from 
44 to 56 inches, according to the blast pressure, method and speed 
of working, concentration to be effected, and so forth. The length 
in many cases is between 15 and 25 feet, when the furnace may be 
conveniently worked in connection with one large settler. The 
capacity of such a unit naturally depends on the conditions of working ; 
it may be taken roughly as from 4 to 6 tons of material per square 
foot of hearth area per twenty -four hours. 

Foun/lations. — The furnaces are built upon a foundation which is 
essarily very strong, being usually either of solid rock or of concrete. 



Bottom Plate. — The bottom plate of the furnace usually carries 
part of the weight of the lower tier of water-jackets as well as the 
furnace burden, and is supported, some distance above the ground, on 
screw-jacks leaving an air-space below the furnace, which allows of 
convenient access for repairs or adjustment. The height of the 
construction is thus raised to a convenient distance for adjustment 
to the discharge to the settlers. The bottom plate should consist of 
sectionised water-cooled cast-iron plates bolted together, Avith a thin 
layer of brickwork placed above, to protect them from the corrosive 
influences to which they are subject. There is a slight slope towards 
the slag-notch. The actual working bed of the furnace is however, a 
chilled crust of material which sets on this bottom owing to radiation 







t ■ I L 



Tapping Breast- Cast Iron 
2 Reci. 

Fig. 39. — Tapping Breast of Blast Furnace, Cananea (seep. 139). 

below, and which, when suitable precautions have been taken, 
usually adjusts itself naturally whilst the furnace is in operation, by 
what may be termed automatic radiation. Thus, apart from the \\'ater- 
cooling devices, if the working bottom wears down towards the metal 
plates, the loss of heat by radiation through the thin layer of material 
causes a chilling effect which leads to a thickening of the crust. 
Should the crust thicken unduly and so threaten to interfere with 
the discharge, the radiation is decreased owing to the thickness ; 
and the high temperature which prevails upon this layer causes a partial 
melting so that it gradually becomes thinner again — thus regulating 
itself for the most part automatically. 

Water- Jackets. — The usual height of the modern furnace, as reckoned 



from tap-hole to charge floor, is roughly from 14 or 16 feet up to 20 feet, 
water- jacketed all the way. The sides and ends of the furnace are 
constructed of sectionised water-jackets arranged horizontally in tiers 
and vertically in panels. There are usually two, occasionally three, 
tiers, suitably stayed and supported. The practice as regards the 

Fig. 40.— Riveted Steel Water-.Tacket, showing Tuyere Holes and Water Inlets, etc. 

(P. & M. M. Co.). 

shape and arrangement of the jackets varies greatly. It was formerly 

I not uncommon to work with three tiers of jackets for the sides ; of 
Wiese the lower tier extended only from the sole-plate to the level of 



varying from 2 feet 6 inches to 4 feet. These were most used when 
the discharge to the settler was situated at the side wall of the furnace. 
Above these jackets was situated the second tier through which the 
tuyeres passed ; these build up the boshes of the furnace, and are termed 


Fig. 41. — Transverse Section through Modern Blast Furnace, showing Arrangements of 
Boshed Lower Jackets, Upper Jackets and Plates, Stays and Supports, etc. 

the '* bosh '' or '' tuyere '' jackets. In most modern furnaces these 
two tiers of lower jackets are replaced by one set of panels of from 7 to! 

[To face page 138. 

42. — Interior of Anaconda Blast Furnace, Bhowing Jacketing, Tuyere Holes, 

and Bridge. 


10 feet in height, the jackets being given a shght slope towards each 
other at the bottom, so as to form a very small bosh angle ; the con- 
traction is about 8 inches. This improvement does away with a good 
deal of the jointing otherwise necessary near the hottest parts of the 
furnace, and thus lessens the danger of leakage at these points. 
The water-cooled breast-plate containing the opening for the escape 
of the products is now put in position as a separate piece, well secured 
to the rest of the jacketing (Fig. 39). Above the lower tier of jackets 
is placed the upper series, often from 7 to 9 feet in height, Avhich 
carries the a^ alls of the furnace up to within a few feet of the charging 
platform. These jackets are parallel, and no bosh is given (see Fig. 41). 

The end jackets are usually built in two tiers only, the upper, 
7 feet to 7 feet 6 inches, as a rule, and the lower, 8 feet to 9 feet 6 inches, 
according to circumstances ; in the smaller furnaces the end wall may 
sometimes consist of a single jacket only. They are vertical, no end 
bosh being allowed. The end jackets are each single panels, whilst 
the side walls are built up in panel sections, the width of which vary, 
but are often 7 feet to 7 feet 6 inches wide, the panels being bolted or 
clamped together and strongly stayed. 

The water-jackets are constructed of flanged steel plate, the inner 
sides of which are w to | inch thick, the outside I to ^ inch. The 
seams are flanged outwards, so as to prevent joints, etc., being 
exposed to the inside of the furnace. The water space between the 
two plates of the jacket is from 3 to 4 inches. 

It is usual to support the weight of these jackets on I-beams carried 
by the upright columns ; very strong bracing and tieing is also neces- 
sary in order to prevent the side walls from bulging by the great 
pressure to which they are subjected. In order to protect the 
jackets themselves from buckhng by the forces acting upon them, 
they are strengthened inside the water space by a series of X bands, 
which run vertically downwards between the plates, and are rivetted 
to the outer side — this device is found not to interfere unduly with 
the proper circulation of the water. liCakage between the joints of 
the separate jackets is prevented by asbestos packing. In spite of the 
strong binding and bracing of the walls in this manner, the connections 
so devised as to allow of their being unfastened very easily, so 
that jackets may be readily disconnected and taken down when it 

jomes necessary to do so. 

Arrangements for the water supply to the jackets vary consider- 
ibly. In locahties where a plentiful supply is available, each jacket 
its independent outlet and inlet pipes ; in other cases it is common 

arrange an independent feed to each set of panels, water being 


supplied first to the jackets of the lower tier, and being discharged 
from them to the jackets situated above. The supply pipes for the 
various jackets branch from water main pipes running at the sides 
of the furnace. 

The tuyere or bosh jackets are pierced horizontally at intervals 
of about 1 foot, with a line of 5- to 7-inch holes for the fitting in of 
the tuyere pieces. These are formed of steel thimbles, of |-inch metal, 
which haA^e a slight taper, fitting secured against the inner plate and 
ri vetted to the outer one, thus allowing of read}^ replacement when 
necessary (see also Fig. 40). Above the side jackets of the furnace 
there is usually a heavy mantel-plate, 2 feet to 2 feet .6 inches high, 
with a sloping front, and surmounting this are apron plates, 1 foot 
6 inches to 2 feet high, inclined at 45"^, constituting a hopper which 
directs the charge towards the centre of the furnace in such a way 
as to keep the fines nearer to the middle line, and thus leave the 
sides of the charge more open, in order to ensure more regular 

Superstructure. — The jacketing, together with the apron and mantel 
plates carry the structure up to the charging floor. Above this is the 
superstructure with the arrangements for taking ofl" the furnace gases, 
and for the feeding of material for the charge. In many cases the 
general practice still prevails of constructing the walls of this portion 
of brickwork, often about 14 feet high, surmounting this with a hood 
of metal from the top or sides of which large ofT-takes carry the 
furnace gases to the dust chambers, and thence to the flue system and 
stack. Modifications in the design of the blast-furnace superstructure 
have been, however, in course of progress at many works, particularly in , 
connection with the employment of automatic or mechanical charging 
appliances and the taking-off of the gases below the feed-floor level. 
This is specially the case at plants operating the pyritic process 
and where the gases are to be utilised for acid manufacture, as well I 
as in connection with the treatment of smelter fume. Several furnaces 
are also at work using either metallic water-cooled or air-cooled tops, 
from which the removal of accretions is stated to be very readily 

Some of the most recent developments in the design of blast- 
furnace superstructure have been described by Emmons in reviewing 
the experiments at the Copperhill Smelter, Tennessee. The gases here , 
are used for acid-making, and are sent to Glover towers under some 
pressure. The furnace top consists of cast-iron corner-posts and 
dividers, the walls and ends laid up with brickwork, surmounted 
by a tubular top of the Shelby type from which the gas off -takes lead. 

[To face page 140. 

Fig. 43. — Showing Upper Jackets, Apron and Mantel Plates, and Superstructure ol 
Blast Furnace, Anaconda. 



The horizontally pivotted doors open inwards and fit tightly. These 
arrangements are stated to be very satisfactory. 

The charging platform, suitably supported on vertical columns, 
runs at the upper level, being provided, on either side of the furnace, 
with tracks of rails for the charge cars. The charging doors usually 
correspond in position to the panels of water-jackets, and are situated 
along the whole length of each side furnace-wall, the bottom of the 
charging opening being flush with the floor. They are generally moved 
up and down in the grooved guides of the upright columns between 
them, and are of sheet steel suitably strengthened, from 6 to 7 feet 
wide and 4 feet 6 inches to about 5 feet high, supported by wire-rope 
and chains, and operated by compressed air cylinders. 

The Air Supply to the Blast Furnace. — The quantity of air required 
by the blast furnace varies very widely with the class of work, rapidity 
of output, character of charge, and general smelting conditions. It 
may be stated roughly as being from 300 to 500 cubic feet of air per 
minute per square foot of hearth area, at a pressure of about 40 to 
50 ozs. per square inch. 

The rotary blower of the Roots or Connersville type is very well 
suited for the supply of these enormous quantities of air at moderate 
pressures, but for blast at higher pressures the air leakage becomes 
excessive, and piston-driven blowing engines become almost a necessity. 
Such improvements have, however, been made in rotary-blowing 
appliances within recent years that most blast-furnace plants are 
equipped with blowers of the rotary type, which are found highly 
satisfactory. The air is brought along blast mains of considerable 
size — about 30 inches diameter — to the furnace building, thence to the 
bustle pipes of 24 inches diameter, which surround the furnace, from 
which branch off the pipe connections (5 or 6 inches diameter) for the 
tuyeres. The practice of equipping each furnace with its own blowing 
unit is fairly general, making the necessary reserve connections in case 
of temporary breakdown ; many smelters, however, adopt the system 
of delivering the air from all the engines into one large common air 
main, making the necessary connections from this to each separate 
furnace. The importance of avoiding leakages is recognised, and the 
requisite valves for regulating and controlling the air supply are 
arranged for. 

From the bustle pipe the air passes down the pipe connectionfl 
which are attached by flanged joints, thence to the tuyere pipes, 
which are of cast iron, the blast being regulated by valves. 

e actual form of tuyere employed varies considerably, each smelter 
lly having its own special devices for the convenience of repair, 




renewal, and fixing, as well as for valve regulation and punching. 
The tuyere is held against the face of the jackets by bolts, leakages 
being prevented by asbestos packing. 

The tuyeres are usually 4 J to 5 inches in diameter, and are generally 


1 ^ » 

m: j/:^''m'W€^ 


Fig. 45. —Blast- Furnace Shell, with Air Connections (P. & M. M. Co.). 

Spring Cap 
Cast Iron 36 Req. ni 

^ Blast Connect/Off^ 
<_ Cast Iron -seFeqr- 


Asbestos Gasket 
^ H'et before 

J?* . .. un F, 

Nipple for Tuyere ^^ '^il 

JCReq. CI. Thimble-26ffsq.\y 

ThisRecess cut topreventOre 
striking sharp Edge of Bead. 

Tuyere - 36 Req. C I 
Fig. 46. — Details of Tuyere, Cananea Blast Furnace. 



placed about 12 inches apart. Air is supplied only through the side 
jackets, and not at the ends of the furnace. 

Heating the Air Blast. — The advisability of heating the air-supply 
for copper blast-furnace smelting has been the subject of very con- 
siderable discussion, the question requiring consideration both with 
respect to its influence on the rationale of the smelting operation as 
well as from the economic standpoint. The matter is dealt with more 
fully in connection with pyritic practice, from which point of view 
Peters has reviewed the subject exhaustively. It may be here stated 
that there appears to be no advantage in preheating the air when the 
true pyritic process is operated, and actual trial has resulted in the 
rejection of the method at the smelters practising this work. 

Where, however, coke fuel to any considerable extent is employed 
on the charge, a supply of heated air through the tuyeres may result 
in an increased rapidity of smelting, as well as in the production of 
hotter and more fluid slags. Especially in partial pyritic smelting and 
more particularly when working charges which contain but little 
sulphide and where the employment of much coke is not advantageous, 
the use of preheated blast may be economically very useful. In such 
cases, the heat production in the furnace is not so fundamentally 
bound up with the thermo-chemical reactions of slag formation as it is 
in true pyritic smelting, and therefore the enhanced intensity of com- 
iistion of coke-fuel at the tuyere-zone by the use of hot air may exert 
«*ii important influence in improving the furnace operation and in 
decreasing the amount of coke-fuel required. In many such instances 
indeed it has been chiefly the economic factor with reference to the 
cost of installing and operating suitable devices for warming the air- 
supply which has determined the question of adopting this system. 
\.s is well known, the use of a supply of heated air causes a largely 
increased calorific intensity from the combustion of coke, resulting in 
iiigher temperature at the tuyere-zone, under which circumstances the 
harge materials are smelted more rapidly, and the resulting products 
<»re more fluid, whilst slags of higher silica content (sometimes 
economically advisable) can be conveniently worked with. 

The devices employed for the preheating of the blast vary con- 
siderably — cheapness, capacity, simplicity in design and operation 
eing the main essentials. 

The utilisation of the waste heat from the smelting furnaces or 
products would suggest itself as an economical method for accom- 
plishing the warming of the blast, but in practice several difficulties 
are encountered in efficiently making use of this heat. Heat is avail- 
able from two sources, either from the furnace gases or from the 


hot slag. The very successful operation in cast-iron smelting, of 
hot-blast stoves worked by the " waste gases/' cannot, however, be 
applied to copper blast-furnace smelting, since the gases in this case 
do not possess similar calorific value owing to the small proportions 
of carbon monoxide present. Further, the temperature of these 
gases is not sufficiently high to allow of the effective application of 
the regenerative principle using brickwork chambers. In consequence, 
the use of metal pipe-stoves offers the only method of utilising the 
heating values of the furnace gases, but their comparatively low 
temperature does not afford sufficient heat for the warming of the 
large quantities of air which are required at the tuyeres. 

The much higher temperature of the reverberatory furnace gases 
offers, however, much greater scope for their utilisation in this respect, 
if both classes of furnace happen to be in operation at the plant and 
if they are conveniently situated for the purpose. 

At several smelters, blast furnaces have been equipped with hot- 
blast "tops'' for the purpose of preheating the air supply, the air- 
heating pipes being exposed to the gases in the upper portions of the 
furnace. The Giroux blast-heating device has been installed on 
furnaces at smelters in Mexico and Arizona, whilst at others in the 
same localities, the Mitchell system of baffle passages has been 
successfully used. The Kiddie system of running the blast pipes 
through the dust chambers has been tried at Tyee, B.C. The advantages 
of thus utilising the heat of waste gases have generally, however, been 
found to be more than balanced by the extra costs involved. 

Efforts have been made to use the heat contained in molten slag 
for warming the air, but owing to the low conducting power of these 
materials, and the difficulty of bringing extended surfaces in close 
contact, the method has not proved itself very efficient. Blast is 
occasionally warmed by passing the air through tunnels in which bogies 
of molten slag are allowed to remain for some time. 

When methods of utilising waste heat from the furnace products 
fail, the fuel-heated iron pipe -stove is generally employed. Since the 
temperatures required are comparatively low, and the margin of 
profit involved by the use of hot blast is usually small, the use of the 
cheapest class of fuel available is imperative ; but many classes of fuel 
unsuitable for other purposes may find useful application for this 
work. ^ ^ 

The stove is of the usual U L cast-iron pipe form, designed to 
give the maximum exposing surface, suitably strengthened and 
protected from direct action of the fire. Much valuable information 
on the advantages, disadvantages, and appliances for blast heating 



was afforded by the smeltermen who contributed to the symposium 
on " Pyrite Smelting/' Avhich Rickard edited for the Engineering and 
Mining Journal. 


Mathewson, E. P., " The Development of the Modem Blast Furnace." Eng. and Min. Joum., 

May27, 1911,p. 1057. 
Wright, Lewis T., ''Metal Losses in Copper Slags." Bulletin Amer. hist. Min. Eng., 1909, 

Sept., No. 33, p. 817. 
Shelby, Geo. F., "Cananea Blast FuiMaces." Engineering and Mining Journal^ April 25th, 

Emmons, N. H., "Copper Blast-Furnace Tops." Bulletin Amer. Inst. Min. Eng., Feb., 

1911, p. 119. 
'•Heating Blast." Engineeriwj and Mining Journal, June 16 and Sept. 15 and 29, 190(i. 
"Pyrite Smelting," T. A. Rickard. 
Also the Authors already referred to, Austin (p. 80), Gowland (p. 17), Peters (p. 80). 




Chargre Calculations — Chargringr — Workingr — Disposal of 
Products— Pyritic Smelting"— Sulphuric Acid manufac- 
ture from Smelter Gases. 

Charge Calculations, — Modern practice aims at the production of 
a matte of converter grade, containing usually from 40 to 50 per 
cent, of copper, and preferably in a single smelting operation ; except 
in true pyritic work.* 

Full analysis of the whole supply of material available at the 
smelter is essential, as well as a report on the quantities of each separate 

The first step in the charge-calculation is the computation of the 
total weights of copper, iron, and sulphur available for the smelting 
campaign ; from these quantities the losses of copper and sulphur to 
be allowed for during the operation itself, as based on previous 
experience, are deducted. The balance indicates the quantities of 
these elements from which the matte and slag can be produced. The 
copper is transformed into matte, in which product it may be 
regarded as existing in the form of copper sulphide, Cu^S, and the 
sulphur required for this combination with the copper is calculated 
from the relation — 

CugS = Cu2 : S :: 2 x 63-5 
:: 127 
:: 4 

1 approximately 

Thus every unit of copper combines with one-quarter of its own 
weight of sulphur. 

A matte of converter grade containing, say, 44 per cent, of copper 
is constituted as follows : — Copper, 44 per cent. + sulphur, 1 1 per 
cent., or copper sulphide, 55 per cent., the remaining portion of the 
matte being iron sulphide, which amounts to 100 — 55, or 45 per cent. 

* This point will be dealt with in due course. 


Assuming as a first approximation that this iron sulphide has 
the formula FeS,* the proportions of iron to sulphur in this material 

are Fe : S :: 56 : 32 

:: 7 : 4 

hence yV of the remaining 45 per cent, of the matte is iron and yV is 
sulphur — that is, the matte contains in addition, iron 28 parts, 
sulphur 17 parts. Hence the composition of the converter matte 
is approximately — Copper 44 parts, iron 28 parts, and sulphur 
11 -^ 17 = 28 parts. 

The amount of copper for the matte is fixed by the available ore 
supply ; the quantity of sulphur is controlled by the furnace operation 
and charges, as judged from previous experience — the oxidation being 
so regulated that the proper grade of matte is produced. The iron 
required for the matte is next considered. Every 44 parts of copper 
require 28 parts of iron for the production of a matte of the correct 
grade. If the quantity of iron in the materials available at the stock- 
bins be not sufiicient to furnish the amount required, as just calculated, 
ferruginous material must be added as flux, if, on the other hand, 
there is a superabundance of iron available in the charges for this 
purpose, the excess must be fluxed off. 

In this manner the amounts of the constituents for the matte 

production are determined, and the composition and making up of 

the slag-forming constituents are next considered. In this connection 

the local conditions with respect to proximity and cost of suitable 

flux, as well as experience with the previous working of the furnace 

snd ore charges are important factors in determining the type and 

imposition of the slag to be made, whilst in true pyritic practice 

le special conditions of working fix certain limits to the composition 

I the slag, as will be indicated later — the pyritic furnace ** tending 

to make its own slag.'' 

In partial pyritic smelting, the coke allowance and the furnace 

conditions allow of fairly wide latitude in making up the charges for the 

roduction of suitable slags with which the furnace can deal efficiently, 

nee the heat production is not dependent on the formation of any 

j 'articular slag. It is always possible to add extra coke for the 

purpose of melting the slag desired. 

The scientific principle governing the calculations for slag composi- 
lon is the proper proportioning of acid and basic constituents. This is 
ased upon the oxygen ratio — i.e., the proportion of oxygen in the 

* The iron sulphide of a copper matte in probably the eutectic of the iron-sulphide : iron 
rie« containing about 85 per cent, of sulphide. 


acid constituents compared with that in the bases. With the doubtful 
exception of akimina in certain cases. siHca constitutes the entire 
acid portion of most copper-smelting slags. 

The requirements for a satisfactory slag are that it shall be — 

(a) Fusible at the temperature of furnace working. 
(h) Fluid and run easily. 
(c) Of sufficiently low specific gravity as will allow of good settling 
and separation of the matte or metallic products. 

It is well known that within certain broad limits of silica content, 
slags will fulfil these conditions to a greater or less extent, whilst the 
most suitable and economic slag under any particular circumstances 
is decided, as stated above, by the composition of the charge, the 
quantity and character of the available fluxes, and the previous 
experience with the furnace. The limits of the silica content for 
suitable slags as just indicated are fixed by several well-known general 
properties of the silicates. 

Speaking broadly, and from the point of view of the more or less 
ferruginous silicates constituting copper-smelting slags, the more basic 
silicates — such as the subsilicate class (oxygen in acid : oxygen in 
base < 1 : 1) — are generally, characterised by high formation-tempera- 
ture, and by being very fluid, thin and fiery, dense and corrosive. 
On the other hand, the more acid siHcates, such as those of the multi- 
silicate class (oxygen in acid : oxygen in base > 2 : 1) are characterised 
by lower formation-temperature and low density, and by being thick 
and viscous. 

As the silica content within this range of silicates increases, the 
melting point is lowered and the specific gravity is reduced, features 
which are very advantageous from the point of view of the production 
of clean slags. Their fluidity, however, decreases, and a very high 
temperature is thus required in order to render them sufficiently limpid 
to run freely from the furnace. On this account the highest proportions 
of silica usually considered feasible in a slag, correspond to the bisilicates ' 
of the representative composition, MO . SiOo. With high temperature : 
conditions in the furnace and rapid working, such slags can be dealt 
with successfully, and if the charges are necessarily highly siliceous, it ; 
may be advantageous from the economic point of view to work with ' 
this class of slag. j 

In proportion as the silica content gradually decreases and as they 
become more basic, the silicates are more and more corrosive and 
fiery, and especially in the case of the iron silicates, they gradually 
attain such a high specific gravity that efficient settling of the matte 


is not possible. In addition, the more basic the sihcate the greater 
is its dissohing power for sulphides, hence high copper losses in the 
slags result from these combined causes. Such basic silicates possess, 
however, the advantage of marked liquidity, and of flowing from 
the furnace in a thin Hmpid stream. The high density and the solvent 
power of basic slags thus fix a limit to the composition which is 
considered economically suitable, and the lowest proportions of silica 
usually worked with correspond to the monosilicates represented by 

I the formula 2M0 . Si02. Slags containing a greater proportion of base 
(usually iron) possess too high a density to permit of clean settling. 
Tn practice, therefore, the majority of slags are mixed silicates of a 
1. composition ranging between the limpid but somewhat dense mono- 
f sihcate and the hghter but more viscous bisilicate, correspondinj:j to 
sihca contents of from 30 to 48 per cent, of silica, and within the 
limits of 35 to 45 per cent, of sihca most copper blast-furnace slags will 
be found. The composition roughly corresponds in a large number 
of cases to that of the sesqui-silicates of the general formula 
4M0 . SSiOo (oxygen in base : oxygen in acid :: 4 : 6 :: 1 : I J). 

As is well known, mixed silicates — i.e., silicates of two or more 
bases — are generally characterised by the properties of increased 
fusibility, and often of increased fluidity, and their employment is 
usual and generally advantageous in smelting practice. The relative 
proportion between the various bases in such mixed silicates is 
largely a matter depending upon the prevailing conditions at the 

In modern smelting, particularly where partial pyritic work is 
conducted, and where fairly siliceous charges are worked, a slag running 
about 40 per cent. Si02 is aimed for, iron and earth oxides constituting 
the remaining 60 per cent, or so. In cases where this quantity of iron 
is present in the charge, the slag may be constituted chiefly of iron 
sihcat'C, but even in such instances the advantages of lime additions 
are marked. When iron is not available in sufficient quantity, the 
extra fuel costs and working difficulties of running with more 
siliceous slags would render their production undesirable, and the 
purchase of limestone or similar earthy flux is particularly advan- 
tageous. The purely iron sihcates are usually dense, and thus tend 
^^to hold up copper values both in mechanical suspension as well as in 
^Molution ; the addition of lime, which has a marked efl^ect in reducing 
^Khe specific gravity, permits of more basic slags being worked with, 
^Hrhere necessary, without such heavy losses in the slag. 
^H The presence of hme silicate with the iron silicates has a marked 



siliceous, whilst on account of the lower atomic weight of calcium, 
lime will, weight for weight, flux off a greater quantity of silica 
than will ferrous oxide. In forming a slag of similar oxygen ratio, 
thus — 

Mono-silicate of lime, 2CaO . SiOg, Lime : silica :: 112 to 60, or 1 part to 0-54 part. 
Mono-silicate of iron, 2reO . SiOo, Iron oxide : silica :: 144 to 60, or 1 part to 0-42 part ; 

hence for the production of a slag of the same oxygen ratio, less 
weight of lime would be required to flux off the same weight of silica ; 
in other words, the replacing values of the two oxides are as 112 to 
144, or 7 to 9. 

Of the other bases which are occasionally present in slags, the 
proportions of the oxides of magnesium and zinc are sometimes con- 
siderable, the calculations being analogous to the previous cases. 
The case of alumina is anomalous, and its behaviour in slag production 
is not definitely understood. Many experienced workers hold the view 
that it tends to act either as acid or base, according to the proportions 
of silica. Thus, in a very siliceous slag, alumina in moderate quantity 
behaves as a basic oxide, forming aluminium silicates, and in very 
basic or low silica slags the alumina appears either to neutralise some 
of the excess base, acting as an acidic oxide, or to dissolve as such 
in the slag, whilst in intermediate cases it possibly behaves partly 
as an acid and partly as base. This view has recently been questioned, 
and it has been suggested by Shelby that alumina always acts as an 
acid in the formation of slags. The matter is thus one which requires 
further considerable investigation. 

Usually neither alumina nor zinc oxide behave very satisfactorily 
in the furnace when present in large quantities, tending to thicken 
the slags and to promote viscosity. 

Anaconda Practice in Charge Calculations. — An example of 
some of the practical considerations which enter into the calculation 
and making up of charges is well illustrated in certain particulars of 
the practice as conducted at Anaconda. Details of the materials 
charged over a period of one month are indicated in Table X. 
The important charge constituents available in large quantity 
include : — 

Per cent. 


Per cent. 

Per cent. 

Per cent 

First-class smelting ore. 

. 8-6 




Concentrates, . 

. 10-9 





. 5-0 




Lime-rock (flux), 

r\i J J. _i ^1 • 



Old converter slags and residues. 




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The other constituents used in the charge comprise varying quan^ 
tities of materials which accumulate round the works, and which, 
being rich in copper values, it becomes useful and essential to clean 
up. For the calculating of the furnace charges, the amounts of cupri- 
ferous material available at the stock-bins are reported to the blast- 
furnace department. The quantities decided upon are divided among 
the number of charges which are considered likely to be a\ orked off 
during the day, this number averaging about 1,100. The result of this 
calculation indicates the amount of each kind of material to be 
weighed for the separate charges ; the analysis of each constituent being 
naturally known. The materials available for smelting are highly 
siliceous in character, the first-class smelting ore, of which large 
quantities are treated, giving a strongly acid composition to the 
charge ; copper-bearing basic materials suitable for fluxing are not 
available in large quantity, and this necessitates the purchase of 
barren lime-rock, this item being the largest of the blast-furnace 
charge. In making up the charge sheet, as large a quantity of con- 
centrate as possible is included, since this constituent is not only high 
in copper values, but owing to a high iron and sulphur proportion, 
it increases the fuel value of the charge, the influence on the coke 
consumption being very marked. The concentrate further forms 
a base for the matte, and introduces iron, of which there is a 
shortage, into the slag, thus reducing its too-siliceous character and 
lessening the quantity of Hme which it would otherwise be necessary to 
procure for the purpose. 

The briquettes are next worked in to as great an extent at possible, 
since by this means the large stocks of settling-pond slime and of 
screened fines are reduced and their 5 per cent, of copper is extracted. 
The whole stock of old slags and residues is used up on the charge, 
these materials introducing considerable amounts of copper, • whilst 
being irony, they further help to reduce the acidity of the slag, thus 
saving the employment of the lime-rock otherwise required for fluxing. 
The total quantity of copper, iron, and sulphur available being then 
calculated, and the allowances for sulphur elimination and for the 
copper loss on smelting (2 to 7 per cent,), as based upon previous 
experience, being deducted, the amount of iron required to constitute 
the 45 per cent, copper matte is estimated. From this figure the 
FeO remaining for slag production is determined. The silica introduced 
by the above materials is also known, and the amount of lime-rock 
required to produce an easily running slag is next calculated. The 
slag which is found by experience to give the most satisfactory running 
has a composition of about — 


SiOo, . . . . .41 per cent. 

Fed, 19 „ 

CaO, 29 „ 

Variations from this composition, especially as regards higher 
silica contents, immediately introduce difficulties, increasing the 
expense of furnace rumiing, by requiring more fuel and care in working, 
reducing tonnage, and producing a slag which runs far less freely. So 
that although the large quantity of siliceous material at hand might 
tempt the management to work with a more sihceous slag, and so save 
the procuring of such large amounts of barren lime-rock, the cost 
of this material is much more than compensated for by the advantages 
which result from the working with a slag which contains only about 
40 per cent, of silica. 

The quantities of the charge constituents thus calculated, divided 
by the hkely number of charges to be worked, are entered up on the 
charge sheet, which is handed over to the charge foreman. 

The Charging of the Blast Furnace. — The method of " hand 
charging,'' as employed in the older processes of working, when using 


Fig. 47. — V-Shaped Charging Car, indicating Mechanism for Release and Tilting. 

small furnaces of small output, possessed several theoretical advantages, 
but it is essential in modern practice, where at least 300 tons of charge, 
and often much larger quantities, are fed into the blast furnace daily, 
to employ mechanical means for charging. At many smelters, how- 
ver, the coke is added separately, from barrows. 
Care in the charging is now recognised as being of special 
importance for successful blast-furnace operation, especially for the 
purpose of procuring the correct distribution of coarse and fine 
material. The principle of keeping the sides more open by distributing 
le coarser materials against the jackets and keeping the fine parts 
arer to the centre is often favoured, since this device reduces the 



tendency to crusting by the finer sulphide particles against the walls. 
It is partly with this object in view that the mantel and apron 
plates are arranged in the hopper form, whilst at the same time the 
distance between the top of the charge and the feed-floor level is 
maintained at such a height that this desired distribution of the 
fresh charges is obtained. 

The practice still commonly employed is to feed the materials 
from side -dumping cars (of very varied design) brought along in a 
train drawn by locomotives and travelling along tracks running 
at each side of the furnace. A form of car frequently used has a 
V-section, and it is secured in a vertical position whilst in transit by 
some form of catch-pin device, which is readily released when it is 
required to tilt the car for charging. 

Another form, employed at Anaconda, has a z \ shaped section, 

the sides of which are pivoted and admit of being very readily secured 
or unfastened as desired. The car bottom itself is tilted by connecting 
it with a compressed air lift by means of a hook situated at the side 
of the car remote from the furnace. The material is thus discharged 
along the inclined chute so produced. 

An interesting method is employed at the Granby Smelter, where 
the Hodge car and the end-feeding method are in use. The cars, which 
have a double-hopper discharge, are divided into four compartments 
by vertical plates. These cars enter at the ends of the furnace through 
suitable openings at the level of the feed-floor, and run by small wheels 
on tracks which are built inside the furnace along the side of each 
vertical wall. In this manner a straight vertical fall for the charge 
is arranged, and this affords the best control of proper distribution. 
The furnace holds three cars at a time, and there are patent openers 
and closers for manipulating the end doors of the furnace, as well as 
for releasing the hopper-bottoms of the cars. 

A particularly ingenious and successful device is in use at the 
Ducktown Smelter of the D. S. C. I. Co.,* Tennessee, where the pyritic 
process is operated. Careful charging is here held to be one of the 
great essentials for successful working of the process, especially in the 
narrow furnaces in use, where the dangers of crusting are greatly 
increased. The principle of working is, that by dropping the charges 
vertically downwards, having previously arranged the materials in 
the desired order across the furnace, they will fall into the position, 
and be distributed just as desired. The Freeland charger is a kind of 
conveyor belt made of overlapping steel plates, which is exactly the 
length and width of the furnace, so that when the machine is brought 

* Ducktown Sulphur, Copper, and Iron Compan}'. 



Fig. 48.— End View of Blast Furnace, showing Tilting of Charge Car, Anaconda. 

Locki, Doors Closed 

^^W*<'- - 

End View 

Fig. 49.— Hodge's Charging Car. 


over it, the furnace opening is entirely covered. The conveyor is 
carried on a frame mounted on wheels, and this is moved forward am 
backward by a motor in the front, near which is seated the chargemai 
who is also the motorman. An independent switch and gearing causes 
the belt to move round and thus deposit its charge over the end. 
In front of the frame is a strong catch, fitting into a recess on the 
cover of the furnace, which is water-cooled and mounted on wheels ,1 
so that as the conveyor is brought into position the cover is moved 
back. All these run along a track which extends below the stock feed- 
bins in the same straight line. The furnace gases are drawn off below 
the feed-floor. 

The method of working is to bring the charger under the bins and 
to drop the various materials for the charge — weighing 2 tons — on to 
the belt. By deflectors on the ore chutes, the charge can be directed 
to any desired position across the belt, and material is thus deposited 
near the outer or inner side as desired — in falling into the furnace it 
is found to take the same position that it had on the plates. The 
charger moves forward and reaches the furnace top, the catch is 
fastened, and as the charger now advances the cover is pushed back, 
the conveyor thus taking its place until in its turn it covers the top 
of the furnace. The motion is now reversed, the conveyor gradually 
recedes, bringing the cover along with it ; meantime the chargeman 
has set the belt-conveyor gearing working independently, and the 
belt thus travelling round and over the end pulleys, discharges its 
burden into the furnace. The disposition of the charge along the length 
of the furnace can be altered at will by increasing or reducing the 
speed of the frame. When the conveyor has at last traversed the 
furnace, the cover is in its place — the charger is now disconnected, 
and goes back for a fresh load. The furnaces are charged eight 
times per hour with 2 tons of material. The operations are fascinating 
to observe, and the control over the disposal of the charge is quite 
complete, whilst the conditions for the operator are not exception- 
ally arduous. Many other suitable devices are in use at different 

At the Cananea smelter is operated an ore-bedding system, the 
store-bins feeding the charge down hoppers through which it falls 
directly into the furnace. A similar feeding system is in use at 
Garfield, Utah. ; 

The lay-out of the plant to allow of the most efficient charging is so 
arranged as to locate the stock-bins at a high level, so that ore is fed; 
directly from the discharge chutes into the cars of the charge trains 
which run on tracks underneath, and these tracks are situated at such^ 



a level that the trains are readily and conveniently hauled to the 
charging platforms of the blast furnaces. 

The charge foreman receives from the blast-furnace department 
his charge sheets which inform him of the amounts of the various 
materials to be loaded on to each car — calculated in the manner already 

X Charge Beit ' 

Furnace Top p g^^^? ' 



Furnace Top 




Fig. 50.- BYeelai)d Charging Machine (D. S. C. & I. Co.). 

indicated. Proceeding to the stock-bins, the gates and chutes of 
which are automatically controlled, he sets the scale of the weigh- 
bridge which is situated under each bin to the desired weight. At 

Fig. 51. — Freeland Cliarger— Details. 

the same time an electric-Ught indicator is switched on in front of 
the particular bins from which material is to be withdrawn, thus 
assisting in spotting the cars and checking the weighing-out. The 

frge train is brought along the tracks running underneath the bins, 
i into each car is dumped the correct amount of charge, usually to 



within 50 lbs., with rapidity and ease. The train then passes to the 
furnace building, where the charges are dumped or otherwise emptied 
into the furnace. 

The Coke Allowance, — As has been already indicated, the coke 
allowance depends largely upon the nature of the charges and the 
individual experience at the smelter. The main principle involved 
is to reduce the coke consumption as much as possible by applying 
the pyritic principle to the fullest possible extent, working as much 
sulphide material into the charge as is economically practicable. 

In partial pyritic smelting, where the coke may constitute from 
5 to 10 or 12 per cent, of the total charge, it is usual not to feed it in 
with the rest of the materials from the cars, but to charge it into the 
furnace separately. The charge foreman puts it in just when and 
how he considers it necessary, and he is encouraged to use as little as 
possible, consistent with proper running of the products at the slag 
spout. In pyritic smelting proper, the small amount of coke is fed 
on to the top of the charge-material in the charge-cars. 

Working of the Blast Furnace. — The top of the charge, which is 
usually some 3 to 5 feet below the level of the feed-floor, appears 
fairly uneven, there being a tendency for it to sink along the middle. 
It is moderately hot, showing practically a black heat except where 
red-hot patches near the side appear in positions corresponding to where 
the tuyeres are situated below. There is not very much fume at the 
feed-floor level if the chimney draft be good, nor excessive agitation at 
the top, unless much fine material is being worked. Sulphide fines 
tend to the formation of accretions near the top of the charge and 
occasionally lower down, also to a considerable extent against the 
walls of the brick superstructure — this is said to be lessened con- 
siderably by the use of water- jacketing at these parts, which also 
greatly assists the barring down of the masses. 

A considerable amount of barring is sometimes necessary when 
much fine concentrate is worked, otherwise a well-managed furnace 
runs smoothly and satisfactorily under favourable conditions. Trouble 
may arise occasionally by leakages occurring in the jackets or spouts, 
but by the modern methods of sectional construction and by the devices 
for time-saving in making the necessary connections, w^orking is usually 
not seriously interfered with for a very long period. Even for the 
removal or replacement of a slag spout, the slag-hole is plugged, and 
the repair is completed within an hour and a-half , by w^hich time slag 
is again running freely over the replaced slag spout. 

The tuyeres are punched regularly two or three times per shift, 
and a steady stream of material issues from the slag notch and over 
the spout to the settlers. 



Disposal of the Furnace Products.— Under ordinary circumstances, 
the products resulting from the blast-furnace operations include — 

(a) The Hquid matte and slag mixture which is given opportunities 
to settle and separate into valuable matte and waste clean slag. 

(b) The '•' gaseous '' products carrying considerable quantities of 
fume and dust which are settled and separated in dust catchers and 
flues, where the solid matter is collected. 

The Matte and Slag. — In modern practice, as already indicated, 
the fluid products of the blast furnace are run out of the furnace as 
rapidly as possible, and flow continuously, as they are formed, through 
a trapped slag notch. So important has this principle of rapid removal 


Slag Spout 

Fig, 52. — Slag Spout, showing Method of Trapping Blast, also Replaceable 
Nose-piece of Spout (A). 

the fluid products become, that the hearth or crucible portion is 
i>^*ing made smaller and smaller. The slag notch, is, in addition, 
placed so low that only so much molten material remains in the furnace 
bottom as is necessary for the regulation of the temperature for 

lintaining perfect fluidity of the materials during their discharge, 
and for avoiding crust formation on the hearth. The depth of 
material remaining in the bottom — that is, the distance from the 
hearth bottom to the slag notch — is from about 8 to 12 inches, 
depending on the conditions just indicated. 

The discharge of the furnace products takes place through the 


trapped slag notch of the furnace, an opening constructed in th( 
tapping-breast or tap-jacket, which is usually a small special jacket 
portion constructed and kept in position separately on account of the 
great local wear at this point (see Fig. 39). The trapping devicd 
is an important and essential feature in connection with the modern 
practice of rapid and continuous running, the principle being to arrange 
a sufficient height of molten material at the outer side of the slag 
opening to overcome the inside blast pressure, and thus prevent the; 
escape of blast with its attendant inconveniences and danger. The, 
flow of liquid material can thus proceed quietly and uninterruptedly, 
The blast is trapped by the construction of a dam in the form of a 
slag spout around the slag opening, of such a shape and secured t() 
the tap- jacket in such a manner and position, that the molten material 
before overflowing at the end, fills the spout and thus covers the 
discharge outlet of the furnace, trapping the blast so that as fasi 
as the molten products form, a constant stream overflows into the 
settlers (see Fig. 52). | 

The slag spouts are often of sheet steel, sometimes of copper or oi 
bronze, and are from 3 feet 6 inches to 5 feet in length, being separately 
water-cooled units. The discharge at the end is from 12 to 18 inches 
higher than the centre of the slag notch in the tap- jacket through 
which the molten material issues from the furnace. The spout is 
secured to the tap-jacket, being arranged so as to admit of ready 
replacement where necessary. Usually it is bolted to the jacket and 
is securely wedged up against it, being supported at the discharge 
end by the wall of the settler, and the joints are made perfectly tight 
by very careful asbestos packing and claying. The spout lasts for 
several months, the greatest wear being at the end over which the 
molten stream issues, but the hfe has been considerably lengthened, 
with greatly increased convenience of furnace working, by providing 
the spouts with separate easily replacable water-cooled nose-pieces 
of cast-iron which are bolted to the ends, thus taking up most of the 
wear and tear, and allowing of a very ready removal and replacement 
without disturbing the slag-spout connections to the furnace itself. 
These are indicated in Figures 52 (A) and 59. The slag spout is 
protected along its entire length by a hood of clay, by which means 
the stream of matte and slag running down it is maintained hot and 

The position of the outlets from the furnace, connecting to the ' 
settlers, is largely affected by the available floor space and the general 
lay-out and arrangements of the plant. Under suitable conditions, 
and especially with long furnaces, the arrangement of the settler 






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Lug -I Iron Sir 



V 6'-A 

\ Patch Bolts lining 

Fig. 53. — Details of Slag Spout, Cananea. 

Fig. o4.— Slag Spout, showing Method of Support. 



in front of the furnace works very advantageously, leaving the ahgn- 
ment of the blast furnaces free, and allowing plenty room for working 
around the settlers. The settlers are then arranged in the middle line 
of the crucible portion of the furnace, so that working is conducted 
evenly from both ends of the furnace towards the discharge in the 
centre, and the smelting is thus regular and allows of good control. 
At many smelters the discharge of products takes place from spouts 
at the ends of the furnaces, the settlers thus being in alignment with 
them. This plan, under suitable conditions, has several advantages, 
permitting of ready access to the sides of the furnace, even working 
of the furnace by discharge at both ends, and ready co-operation 
between adjoining furnaces and settlers. 

Settlers. — The modern type of settler is often circular in section, 
about 16 to 18 feet in diameter and 5 feet in height, storing about 
40 tons of matte. Other forms, rectangular or oval, are, however, 
also employed. 

The outer shell is of |-inch steel plate bound together by band- 
bolts, the lining is often 9 to 15 inches in thickness, with an inside 
layer of looser stuff. The lining material employed varies greatly, 
according to the grade of matte, character of slag, and working con- 
ditions. The wearing out of the lining depends very largely on the 
class of material passing through the settler, the most rapid wear 
being occasioned by the fiery and corrosive low-grade mattes and basic 
slags, whilst high-grade mattes and more siliceous slags give little 
trouble in this connection. The more corrosive the products, the 
more refractory and hard-wearing must be the lining, and consequently 
the materials employed for the purpose range from chromite, silica- 
brick and firebrick down to loam, according to the requirements ; the 
chief duty is that of being non- corrodible and of protecting the outer 
shell. It is not an uncommon practice to thicken the walls close to the 
tap-holes, where they are subjected to most wear, and often chromite 
is used at these points owing to its power of withstanding the forces of 
erosion. On the other hand, at the Copperhill Smelter of the Tennessee 
Copper Company the settlers have been found to give as satisfactory 
service on fairly low-tenor matte, when lined throughout with good fire- 
brick as with the more expensive materials formerly used, whilst stil 
more recently, sihceous copper ores have been successfully employed 
as lining material instead of bricks. 

There is usually a spray of water from a circular pipe which sur- 
rounds the settler near the top — this playing against the steel sides 
keeps the outside cool and protects the lining. The settler is roofed 
over with slag, except at the back where the stream of matte and 



Fig. 55.— General View of Settler (T. E. Co.). 

Fig. 56. — Method of Lining Settler, Cananea. 



slag enters, and also at those points where the slag overflows. Thel 
slag escapes over short launders attached to the top of the steel casing. 
The position of these discharges depends largely on the arrangement of 
tracks, size of furnace, temperature of working, and quality of pro- 
ducts. Under modern conditions of high temperature and rapid 
working, they are situated as far away from the entrance as possible, 
thus giving fuller opportunities for very quiet settling in a large pool 

Discharge Spout from Furnace ^.^^^T^r'i^^^^ Top 

Matte Launder. 
Fig. 57.— Arrangement for Matte and Slag Discharge from Settlers (T. C. C ). 

and affording gentle overflow of slag with little abrasive action on 
the linings. These outlets may be situated opposite to the entrance 
or at the sides. The discharge spouts for slag may be one or two 
in number, usually of cast iron coated with thin clay, and often roughly 
hooded over with clay. They have replaceable cast-iron nose-pieces 
to faciUtate repair after wearing down. The continual gentle stream 



I of slag runs along launders, where it is either discharged into slag 
bogies and dumped, or much better, is met by a strong stream of 
water ^^ hich immediately granulates it, and washes it along flumes to 
the dumps. 


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I Hole 

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Iron Centre for Tap-holo 

Copper Tap-hole Plate. 

Fig. 58. — Tap-hole Casting and Detail for Settlers. 


The matte tap-holes are generally two in number, situated close 
the bottom of the settler, and usually at an angle of 120° from each^ 
other and from the entrance spout. 

The hole through the brick wall for tapping is about IJ inches in 
diameter, and the matte is discharged through a tapping piece of' 
cast-iron, 6 inches in diameter and 3 inches thick, perforated by a 
1-inch hole. This iron disc has, cast around it, a copper tapping-plate 
about 1 foot in width and 2 feet high, which is recessed into the steel 
sheet of the settler. In the iron tapping-piece is a conical recess, into 
which the conical clay plug is rammed when closing the tapping-hole. 
These iron tapping-pieces withstand the action of converter grade 
matte fairly well, and are conveniently replaced when necessary — j 
about once a month. They are illustrated in Fig. 58. 

The tapping-plate is fixed into position in a special section of the 
shell, known as the launder casting, to which the matte launder is 
secured, whilst a newer form of settler has the tap sections also remov- 
able, so that these can be taken out and the brick renewed during the 
campaign of a furnace, being as readily removable as a furnace jacket. 
The matte launder is of cast iron or of steel, thickly coated with clay 
or suitable material (slime-pond product, etc.) to protect it from 
corrosion. In modern work the steel tapping bar is always rammed 
through the conical plug and tapping-hole until it just reaches the 
matte, so that its withdrawal by ring and wedge is readily performed 
when the matte is to be tapped whilst by this means the tap-hole is 
securely closed. 

The workers are protected from shots of matte, etc., during tapping 
or closing, by means of a slotted sheet-iron hood which can be swung 
back when not required, a convenient and useful as well as necessary 
precautionary device. Matte is tapped from the settlers into ladles 
as required by the converters ; such ladles are constructed of thick 
steel plate, washed with clay, and often lined with a hull of chilled 
material. It is sampled at the runner with each tapping. The tap- 
hole is closed by a clay plug on the end of a dolly which is rammed 
home, and a warm pointed steel bar is then driven through until it 
reaches the matte, being knocked in occasionally as the end is very 
slowly eaten away. Several of the features named in the previous 
sections are well indicated in the photograph (Fig. 59) of the tapping 
platform at the Anaconda Smelter. 

The " Gaseous" Products of the Furnace. — Great variation is to' 
be found in the arrangement at different works for the disposal of the 
gaseous products of the furnace. Reference will be made later to 
the methods employed in connection with pyritic work, and where 



the gases are to be utilised for the production of sulphuric acid. 
Forme rh' the general method, even at the large modern plants, was 
to lead the gases from the top of the superstructure to the off-takes 
and large dust-catcher flues, thence to the stack. 

With the introduction of automatic and mechanical charging 
methods, now being inaugurated to a considerable extent in place of 
dumping from cars alongside the furnace, the method of withdrawing 
the gaseous products just below^ the level of the feed-floor is being 

The off-take flues of the modern furnace are of steel, 4 to 6 feet 

Fig. 60. — Hoppers of Flue-dust Chambers and Tracks for Cars underneath. 

in diameter — lined or unHned according to circumstances — and leading 
to very large dust chambers of varying design, sometimes rectangular, 
often of large circular section, or of balloon- shaped section, etc. In 
all cases these flues are provided with hopper discharge openings at 
suitable intervals, under which cars run on tracks, for the collection 
and conveying of the dust. Arrangements for the further settling and 
oUection of the flue-dust are essential in connection with modern 
last-furnace plants, where blast pressures of from 40 to 50 ozs. per 
inch are employed and where it is often found economical to work 
ith as much fine material as possible, either as such or in an 



agglomerated form ; where too, the dropping of charges from some 
height and the agitation caused by the blast are practically unavoid- 
able. Rarely less than 2 per cent, of flue-dust is made in any modern 
blast furnace, whilst 5 per cent, is by no means uncommon, and even 
larger quantities are often produced. Such dust is, moreover, often 
somewhat higher in copper contents than the original charge, owing to 
the brittleness of copper sulphide minerals, which, being more readily 
broken up, are carried over in the form of fine particles. Hence the 
economic aspect of the recovery of values, in addition to legislative 
requirements, call for efficient collection of these products. 

The gaseous products of the furnace carry solid matter in two 
forms. As a rule, under the usual conditions of copper-smelting 
charges, the larger portion of the solid matter thus carried is in the 
form of very fine particles of charge material itself, mechanically 
suspended and carried over in the current of the escaping gases. This 
is the flue-dust. In addition, values in the form of volatilised metallic 
products are also conveyed by the gases, particularly when lead, 
zinc, arsenic, etc., are present in the furnace charge, and these are 
carried forward in the form of fume. They tend to solidify as the tem- 
perature of the gases becomes lower, although their settling is very 
greatly impeded owing to the exceeding minuteness of their particles 
and also to their dilution ; the problem of separating and collecting 
them is in consequence attended with great difficulty. 

Chambers of enormous capacity are required in order to give the 
fine solid particles an opportunity of settling by decreasing the velocity 
of the gases and by coofing them down, whilst for the settling of fume, 
capacious flues in which are suspended wires or similar devices for 
assisting the process must be adopted. Where large quantities of lead, 
etc., are present some bag-house system of fume filtration is necessary, 
especially if silver be present, since this metal tends to be carried over 
in the leady fume. At the majority of copper smelters such extreme 
refinements are rarely necessary, although modern legislative require- 
ments make severe demands on the managements for the freedom of the 
gases from injurious constituents. 

Dry settling methods and filtration are in general use where such 
separation is required and the use of high-tension electricity has 
been successfully tried at Calif ornian smelters. Wet methods have 
so far not proved economically successful. 

The flue-dust from the flues is dealt with in a number of ways, 
according to the conditions at the smelter. It may be smelted with 
the " roaster-calcines charges '' in the reverberatory furnaces, although 
excessive quantities have proved difficult to deal with in certain 


instances, it may be included in the charges for sintering or briquetting 
processes, and it has been very successfully incorporated with the matte 
in beds when it has been necessary to cast low-grade matte into 
cakes previous to reconcentration in the blast furnace, at a smelter 
employing the pj^ritic process. 

Still more recently, the East Butte Copper Mining Company has 
installed and successfully operated a sintering plant on the Dwight- 
Lloyd principle for the treatment of the flue-dust preparatory to 
blast-furnace smelting. The capacity of the plant is 100 tons per day. 
The material is rendered more or less cohesive by the effects of heat 
alone, but the operation is not yet perfect. (See Mining Journal, 
Jan. 6th, 1912, p. 21.) 

The freed gases finally pass along series of long and capacious 
brick main-flues connecting with all the branch flues, furnished with 
discharge hoppers at intervals, gradually rising and discharging into 
a wide stack of such a height that damage to vegetation in the district 
is entirely prevented. 

Pyritic Smelting. — Modern blast-furnace practice, as has been 
stated, is conducted according to two main systems of working : — 

(a) That in which the heat required in the smelting zone is pro- 
vided by the oxidation of the sulphide materials of the charge — 
Pyritic Smelting. 

{h) That in which coke or other carbonaceous fuel is necessary for 
supplying some of the heat required in the smelting zone of the furnace, 
even when the pyritic effect of the charge is utilised to the fullest 
extent — Partial Pyritic Smelting. 

The term Pyritic Smelting (or pyrite smelting) is thus applied to 
that class of practice in which the whole of the heat required in the 
Milting zone is obtained by the combustion of the ore or matte 
!. iige itself ; it implies the application of the pyritic principle to the 
extreme limit, the use of carbonaceous fuel being reduced to a minimum. 
Ideal working is to feed unroasted ore or matte, together with 
the requisite fluxes, into the blast furnace, and by the action of an 
adequate air blast, to bum out part of the sulphur and iron, the former 
escaping with the furnace gases, the latter being slagged off, whilst 
the copper in the charge is concentrated in the matte product of the 
I^K This type of smelting is conducted at a number of large modern 
l^works, and though up to the present time the use of coke on the 
I^Bliarge has not been entirely eliminated, research and practical ex- 



is not utilised as fuel by combustion in the air blast at the tuyeres^ 
but that it is, in fact, oxidised in another manner at some considerable 
height in the furnace. 

History. — The idea originated with John Holway, of London, who 
sought to extend to the smelting of copper the principles so brilliantly 
apphed by Bessemer to steel manufacture, and who, in a work which 
was published in 1879, suggested and demonstrated the process of 
utilising the heat of oxidation of the iron and sulphur constituents 
of copper-bearing materials for the smelting and extraction of the 
copper. That work is to-day recognised as one of the most masterly 
expositions of the principles underlying pyritic smelting and con- 
verting, and many of the most important and recent developments in 
these branches of work are proceeding on lines forecasted b}'- him. 
Holway 's experiments, conducted on a considerable scale, proved the 
feasibility of the principles underlying the process, which was to 
prepare metallic copper from sulphide ores in one combined series of 
operations in a single furnace unit. Owing, however, to mechanical 
troubles and difficulties of operation, as well as to the ultimate with- 
drawal of financial support, he was unable to carry the process to a 
commercial success, and the single-stage process is at present regarded 
as being beset by almost insuperable difficulties, although the latest 
phases in modern practice are tending towards a realisation of Holway*s 
scheme of working. His paper and published results deserve the 
closest study. 

Inspired by the pamphlet, an EngUsh Company in 1887-8 attempted 
the practice at a smelter at Toston, Montana, and showed the possi- 
bihties of the method, although the plant available did not lend itself 
to completely successful operation. L. S Austin, who took a leading 
part in this work, patented the process in the United States, and 
developed the practice, and in 1891 Dr. Peters conducted a very full 
enquiry into the conditions of working, which placed the system on a 
definite practical basis. From that time the method has developed co- 
incidently with the more empirical practice at many works of replacing 
coke fuel by sulphides to as great an extent as possible. T. A. Rickard 
focussed scientific and practical opinion on the subject in the symposium 
on '' Pyrite Smelting,'' which he called forth and edited, and many 
celebrated smeltermen have contributed to the progress of pyritic 
smelting practice. At the Copperhill Smelter of the Tennessee Copper 
Company and at the Ducktown Sulphur, Copper and Iron Co.'s Smelter 
at Isabella, Tennessee, remarkably good pioneer work was done by ^ 
Parke Channing, Freeland, and others in developing the process. > 
Enormous service has been rendered within recent years by the masterly 


researches and brilliant exposition of Robert Sticht, in which latter 
\^ ork Peters has worthily seconded him. 

Pyritic smelting is at the present time being very successfully 
practised at Mt. Lyell, Tasmania ; at Tennessee ; Tilt Cove, Newfound- 
land : and other districts, w^hilst the smoke problem alone has pre- 
vented for a time a number of other smelters from successfully oper- 
ating the process. 

The Mechanism of the Process, — The mechanism of the changes 
involved in the pyritic process is now fairly well understood in general 
outhne. One of the most important steps in elucidating the matter 
was made by Sticht s discovery that the oxidation area of the furnace 
in pyritic smelting was confined to a narrow zone situated just a little 
higher than the tuyere level ; by actual experiment it was found that 
scarcely any free oxygen existed above this narrow tuyere zone.* It 
thus became evident that the first series of changes near the top of 
the charge were those mainly caused by the effects of heat alone, 
and that only by a second series of changes lower down at the tuyere 
zone were the reactions of rapid and intense combustion and oxidation 
of the sulphides being effected. Finally, at the bottom of the furnace, 
the molten matte and slag collected and ran out. Thus the furnace 
operations proceed in two main stages ; preparation (liquation of the 
sulphides from the charge) in the upper portion, and oxidation and 
fluxing (bessemerising of the liquated sulphides) in the oxidising tuyere 
zone or focus. 

The usual and typical ore charged into the furnace in pyritic 
smelting is impure chalcopyrite (essentially a copper-bearing pyrites, 
Fe8o). When heated in an atmosphere free from oxygen, this pyrites 
loses some of its sulphur and approaches pyrrhotite in composition. 
On further heating in a neutral atmosphere more sulphur is evolved 
and the material approaches FeS in composition, whilst at very high 
temperatures and under favourable circumstances, a still further 
quantity of sulphur is liberated, resulting in the production of the 
well-known fusible iron sulphide, which is the eutectic of the iron : 
iron-sulphide series of alloys, melting at 970° C, and containing about 
85 per cent, of FeS. Thus in the pyritic furnace, free sulphur is 
liberated as such at the upper levels, and passes up the furnace un- 
^^ changed until it meets free air above the surface of the charge, when 
^ht there bums to SO2. The residual fusible sulphide melts, trickles 


♦These views have recently been controverted in an interesting paper by G. A. Guess. 
Notes on Pyritic Smelting," Engineerimj and Mining Journal, 1912, Jan. 13th, p. 113). 
c defines Pyritic Smelting as the production of a Ferrous Silicate Slag from Iron Sulphide 


down, and becomes the true pyritic fuel of the furnace. The copper 
sulphide constituents of the charge are practically unaffected in com- 
position b}^ heat alone, and they pass down the furnace with the rest 
of the charge unchanged until the hotter zones of the furnace are 
reached, when these sulphides also liquate out, become dissolved in 
the melting iron sulphides, and are thus carried down to the oxidising 
zone. Until the sulphides meet free oxygen, no further reactions 
proceed, since they are without action on silica at even the highest 
furnace temperatures. 

When, however, they reach the blast of air which enters the 
furnace at the tuyeres, an intense action proceeds as the sulphides 
become bessemerised. The heat of oxidation of iron sulphide has 
long been known to be very great, and Holway pointed out that 
this heat corresponds to the large quantity of heat which is developed 
by the free roasting of heavy sulphides, compressed into the space of a 
few moments, and thus results in an exceedingly great intensity with 
consequent high temperature. Sulphur is burnt out to SOo, iron is 
converted to the oxide which instantly combines with the white-hot 
silica skeleton that is present and forms an iron-silicate slag, evolving 
still more heat. This slag, with the enriched matte, melt thoroughly at 
the prevailing temperatures, and issue from the slag spout of the furnace. 

The work of Sticht and Peters thus allow of the mechanism of the 
processes being followed during the passage of the materials through 
the furnace. 

At the Mount Lyell Smelter, where Sticht operated, the charge ex- 
tends about 12 feet above the tuyeres. In the upper 6 or 7 feet, 
elemental sulphur is driven off from the pyritic materials by the effects 
of heat alone, and the furnace gases in this zone consist chiefly of 
nitrogen, SOg (from the bessemerising), sulphur vapour, a little CO.,, 
but practically no free oxygen. About half-way down, the temperature 
is sufficiently high to melt out the fusible sulphides from the charge ; ^ 
these liquate and trickle unchanged through the still solid masses of f 
gangue and silica-flux, until they meet with free oxygen of the air 
blast, when they are oxidised and burnt up with great rapidity and 
with the evolution of intense heat. This bessemerising zone extends 
from a short distance above the tuyeres to a point where all the 
oxygen is used up by the iron and sulphur. The distance is variable, 
but is probably some 2 feet or so. At this level the ferrous oxide 
produced is instantaneously seized by the white-hot particles of free 
silica with the production of a silicate slag, the composition of ivhich 
corresponds to the silicate whose formation temperature is eqnul to that 
prevailing in this bessemerising zone. 


Control of the Operations, — It has thus been estabhshed that the 
oxygen of the air blast entering the furnace through the tuyeres is 
practically all expended in this bessemerising of the liquated sul- 
phides in the narrow bessemerising zone, and that it does not operate 
at all by any roasting reactions in the upper part of the furnace, as had 
been formerly supposed. 

From this knowledge it therefore becomes possible to indicate the 
essential factors which control the successful operation of true pyritic 
smelting. The degree of bessemerising depends upon the amount of 
air supplied for the oxidation of the sulphides, and upon the quantity 
of siliceous flux present to slag off the iron oxide produced. 

The actual smelting takes place at the focus where the liquated 
sulphides are instantaneously bessemerised, and the more rapid this 
oxidation, the more intense are the reactions and the higher the tem- 
peratures which result. 

For successful pyritic smelting it is, therefore, essential that there 
shall be present — 

(«) Sufficient sulphides in the charge to give out the heat 

necessary for the smelting and for the thorough fusion of the 

(h) Sufficient oxygen (air) for the rapid and necessary 

oxidation of this sulphur and iron, 
(c) Sufficient free siliceous flux for the satisfactory slagging 

of the iron oxides produced. 

{a) The supply of heat required for the smelting of the charge 
and the thorough fusion of the products depends entirely on the 
intense combustion of the iron and sulphur constituents, and the greater 
the proportion of these materials oxidised per minute, the higher is the 
temperature. As has been already noted, such heat intensity increases 
at a rate greater than the mere arithmetical increase in the fuel pro- 
portion, by reason of well-known thermo-chemical laws regarding 
mass effects. Indirectly, too, the higher the proportions of sulphides 
present, the smaller is the quantity of inert or useless matter which 
requires to be heated and slagged off in the furnace — apart from the 
question of the necessary flux material. Hence the higher the iron and 

I ^—sulphur contents of the ore, the more successfully may true pyritic 
^pBmelting be applied to it. True pyritic smelting may be said to cease 
when carbonaceous fuel requires to be burnt at the tuyere zone in order 

I to supplement the heat derived from the sulphides, and broadly 
speaking, from about 28 per cent, of iron and about 30 per cent, of 



conditions. At Tennessee, with about these proportions, the coke 
consumption on the charge is reduced to about 3 to 4 per cent. ; at Mt. 
Lyell, where the ore runs from 40 per cent, of iron with a corresponding 
quantity of sulphur, the coke consumption amounts to only about 1-25 
per cent. None of this coke probably reaches the bessemerising zone 
at all. 

(h) Being supplied with enough sulphide fuel, the requisite quantity 
of air for the rapid and sufficient combustion of this iron and sulphur 
is essential. The oxygen is used up entirely in the bessemerising 
of the sulphides at the tuyere zone of the furnace, and in consequence, 
not only the heat supply, but also the concentration depends upon the 
amount of oxygen furnished at this point, since the greater the quantity 
of oxygen which is used up, the greater is the amount of sulphur elimi- 
nated and the amount of iron oxidised and slagged off, and in conse- 
quence, the higher is the proportion of copper in the resulting matte. 
In other words, the oxygen supply largely controls the concentration 
effected in the smelting process, and consequently an adequate quantity 
is of the utmost importance. The amount of air theoretically required 
per minute is readily calculated from the estimated capacity of the 
furnace and from the charge analysis. Liberal allowances are required 
for losses, leakages, blower efficiency, etc. ; and the volume necessary at 
the furnace amounts to something like 5,000 cubic feet per minute 
per 100 tons of sulphide. 

(c) Sufficient siliceous flux is required for the satisfactory slagging 
of the iron oxides produced. The presence of the requisite silica on 
the charge is exceedingly important. The iron of the sulphides, upon 
oxidation by the air blast, is converted into iron oxides, primarily FeO. 
This oxide is incapable of existing by itself, but possessing when 
nascent a powerful affinity for silica at high temperatures, it produces 
ferrous silicates, which are, in the main, fusible slag-like products. This 
action is particularly evident in the tuyere zone of the pyrite furnace, 
where the silica is present in a white-hot condition. If sufficient silica 
be not present to combine with the iron oxide produced, the ferrous 
oxide which is exceedingly unstable, finding itself without the neces- 
sary flux, is converted under the continued oxidising effect of the blast 
into higher oxides of iron such as ferric oxide or magnetic oxides, 
materials which are practically infusible, and this results in the produc- 
tion of an infusible sinter which leads to the choking of the furnace. 
On the other hand, if excess of silica be present in the charge, highly 
siliceous and unworkable products result, which will not run out of the 
furnace. Any further excess of silica simply remains unfused and un- 
attacked, and causes the ultimate stoppage of the furnace operations. 



The silica for fluxing is consequently an important factor in con- 
trolling the running of the pyritic furnace, and the provision of the 
requisite quantity, as nearly as possible, is essential, since otherwise 
the presence of adequate sulphide and air blast is not in itself suffi- 
cient to ensure satisfactory working. 

The actual quantity of silica required is determined by the factor 
kno^^-n as the formation temperature of the silicates. Every silicate 
has a definite formation temperature — i.e., a definite mixture of iron 
oxide and silica requires a definite temperature in order that complete 
combination may occur and a chemical compound silicate be formed. 
Conversely, at any definite temperature, only those silicates having a 
corresponding formation temperature to this degree of heat can be pro- 
duced. In consequence, if the oxidation of the sulphides at the tuyere 
zone produces any particular temperature, that particular silicate whose 
formation temperature corresponds to this will tend to be formed, and 
the required quantity of free silica must be present to yield this definite 
silicate with the whole of the iron oxidised. Only a hmited quantity 
of silica can thus be taken up for any definite rate of oxidation of iron 
<^ulphide, and the presence of either more or less silica does not greatly 
tffect the composition of the slag. Thus the concentration (sulphide 
oxidation) is primarily dependent on the oxygen supply, which deter- 
mines how much iron shall be burnt, but the success of the operation 
depends upon the presence of the correct amount of silica to flux off 
this iron oxide. This proportion is fixed by the temperature attained 
at the tuyere zone, which restricts the silicate produced to such a 
composition that its formation temperature coincides with this degree 
of heat. Hence the general law has been deduced and has been 
confirmed in practice, that " a pyritic furnace produces a slag corre- 
sponding in composition to the silicates whose formation temperature 
equals that prevailing at the tuyere zone," accounting for the well- 
known observation " that the pyritic furnace tends to make its own 
-lag." If the smelting operation is to proceed satisfactorily, slag 
approaching this composition will be produced, and assuming the 
air supply to be adequate for the purpose, the absence of the requisite 
.silica on the charge affects the quantity rather than the character 
of the slag. The amount of iron sulphide oxidised depends largely 
upon the presence of silica to combine with the iron oxide produced ; 
HO much will be oxidised as the silica can deal with, and in consequence, 
if the free siHca supply is deficient, a smaller quantity of slag is 

Iormed, whilst the matte will be larger in amount but of lower grade. 
bi addition of silica to the furnace charge under such circum- 


slagging of more iron, and would produce slag of approximately the 
same composition as before, though in larger quantity. 

Deficiency of silica also results in the production of over-fire, owing 
to the fact that the air blast, being unable to bessemerise any more 
iron sulphide at the tuyere zone, passes to the higher portions of the 
furnace and gradually roasts the ore there, thus consuming the 
sulphide fuel of the furnace which might otherwise be most effectively 
used for bessemerising in the tuyere zone. This over-fire, resulting 
from the heat of roasting which is given out in the upper part of the 
furnace, is very disadvantageous in true pyritic smelting, and successful 
control of the process depends on using up the whole of an adequate air 
supply at the bessemerising zone, and on supplying sufficient siliceous 
flux to combine at once with the whole of the iron oxide produced. 
For fluxing purposes it is only the free silica in the charge which is 
effective, since any silica existing as silicate is already in a state of 
combination and thus is not free to act as flux. The combined silica, 
except for its adding to the fusibility of the charge by admixture, 
is very disadvantageous, consuming heat and space, diluting the 
reaction intensities by presenting an inert substance among the active 
constituents, and increasing the quantity of slag which requires to be 

The three requirements — iron sulphide, oxygen supply, and fluxing 
silica — thus bear an intimate relationship to one another in true 
pyritic smelting, and alteration of any one factor requires simultaneous 
adjustment of the others for the production of the same grade of matte 
and slag. The speed and degree of oxidation primarily depend on 
the air supply. The more iron burnt up, the greater is the heat pro- 
duction and the higher the temperature at the tuyere zone, and since 
the more basic slags are known to have the higher formation tem- 
peratures, the basicity of the slags increases with the speed of oxida- 
tion and consequent concentration. 

Ores suitable for true pyritic smelting are not commonly met with 
in practice, and the presence of earthy bases other than iron is not 
desirable. Whilst the advantages of polybasic slags from the point of 
view of reduced formation temperature, increased fusibility and 
liquidity are very marked in ordinary smelting practice, their presence 
is not so advantageous in true pyritic smelting, since they consume 
silica which is required for the iron oxide at the instant of formation, 
and thus tend to decrease the speed of oxidation and concentration. 
Polybasic slags have a lower formation temperature, and in con- 
sequence the production of the highly ferruginous slags of high 
formation temperature which it is desired to make by the oxidation of 


a« much iron as possible is retarded. In addition, the presence of 
other earthy bases in the charge dikites its fuel value ; they may even 
consume valuable heat by requiring decomposition, as in the case of 
carbonates. These considerations are not so important in partial 
pyritic smelting, where the required heat balance can be adjusted by 

The Advantages of Pyritic Smelting. 

(1) The possibility of direct and immediate treatment of highly 
pyritic raw ore in the blast furnace, thus saving all the costs of 
prehminary treatment and handling. 

(2) The saving of the costs of roasting heavy sulphides. 

In former smelting practice, high sulphide contents in a copper 
ore were particularly disadvantageous, since the higher the sulphur 
contents of the charge the lower was the grade of the resulting matte, 
when smelted directly in the blast furnace. In consequence, the 
higher sulphur content necessitated a more complete roasting of the 
ore in order to ensure a high-grade matte on smelting. 

With pyritic smelting the conditions are completely reversed, and 

the charge becomes more suitable for direct furnace treatment as its 

sulphide contents increase, so that the most suitable ores for pyritic 

melting are those in which the greatest saving is effected by their 

not requiring a preliminary roasting operation. 

As has been already indicated, this saving includes labour, plant, 
handhng, time, and interest on capital tied up in the roast yards, 
as well as the avoiding of all the mechanical and other losses connected 
with such preliminary treatment. Thus at Ducktown, Tennessee, 
the material economies effected by the substitution of pyritic smelting 
for the processes involving preliminary roasting amounted to no less 
than 3 to 4 cents per pound of copper produced, in addition to the 
later advantages derived from the recovery of values from the gases, 
and from the improved conditions of life in the district. 

(3) The cost of coke is saved. 

Fuel is one of the main items of expense in blast-furnace smelting, 
and by the substitution of the cost-free natural-sulphide fuel for coke, 
the proportion of the latter required on the charge is reduced from 
^lie 9 to 10 per cent, formerly employed with roasted materials to 
^bout 3 to 5 per cent., and in certain special cases to very much 
laller amounts. 

Diffwulties of the Process.— Tha.t the technical difficulties in applying 

le process on a practical scale are considerable, under present condi- 

)n8 of working, will be understood from the nature of the operations. 

(1) The pyritic process works on a narrow margin of heat, and 



allows of but little flexibility in the conditions of working, since there 
are few factors which can be altered should difficulties in operating 
arise, as compared with the circumstances when a free use of supple- 
mentary carbonaceous fuel may be employed. The only source of | 
heat energy at the smelting zone is in the sulphide charge itself, and 
small variations in the working conditions may readily disturb the 
delicate equilibrium upon which successful working depends. Irregu- 
larities, stoppages, and variations in grade of matte ma}^ therefore 
arise, unless the operations are regulated with exceeding watchfulness. 
In true pyritic smelting the employment of coke for restoring the 
balance or for producing heat required at the tuyere zone is not per- 
missible or practicable, since, as will be indicated later, such coke 
addition would altogether destroy the equilibrium in the process ; 
the grade of matte and the composition of slag would be altered, the 
reactions disturbed, and a restoration to normal pyritic smelting 
conditions rendered almost impossible. 

Difficulties in operation have therefore to be overcome along the 
lines of pyritic action — that is, in the further adjustment and mani- 
pulation of blast, sulphide or siUca supply, or in charging methods, 
etc. — and in practice such careful " doctoring " is resorted to when the 
furnace shows signs of working unsatisfactorily. 

It is very often possible by such careful attention to gradually 
bring a furnace back to smooth running. It occasionally happens, 
however, that the conditions gradually become worse, and the furnace 
commences to show signs of " gobbing.'' This is indicated at the top 
of the charge by the formation of crusts round the side and end walls, 
whilst from the slag spout below, there issues a much reduced quantity 
of thick siliceous slag, together with an abundant stream of thin 
low-grade matte. The furnace gradually ceases running, and it be- 
comes necessary to stop its working, to take down the furnace jackets, 
bar out the debris, and restart operations. This is usually not so 
objectionable a procedure as it might appear, and indeed, within 
certain definite limits, such a course may economically be sound policy. 
In the modern operation of pyritic practice it often pays better to 
risk the occasional gobbing up of a furnace and clear out the debris, 
than to work with so large a quantity of coke as would avoid such a 
necessity. Not only is the modern furnace so designed and constructed 
as to entail but comparatively little trouble in cleaning out in this 
manner, but such practice, even if temporarily a necessary evil, may, 
in places where coke is expensive, and where conditions for pyritic 
smelting are otherwise favourable, be, within certain definite limits, 
actually the most profitable. It is by the taking of these risks, combined 


Avith further experiment and working experience in manipulation, 
such as in charging methods, blast conditions, and the height and 
distribution of charges, etc., that the ultimate continuous and successful 
working at still lower costs may be attained and the true pyritic 
process be worked as continuously as ordinary smelting practice. 
Short campaigns are not, therefore, unusual under the present con- 
ditions of true pyritic smelting, and the cleaning out of the generally 
fairly loose debris is accompUshed with moderate ease, from 24 to 
36 hours being the usual time required to take down, clean out, and 
restart a furnace, whilst the cost of such an operation (chiefly in 
labour) is not, under the circumstances, excessive. At Tennessee, 
hard driving and short campaigns result in lower costs and greater 

(2) The composition of the slag often prevents high concentration. 

It has been indicated that the thermal conditions in the bessemer- 
ising zone of the pyritic furnace tend to the production of highly basic 
slags, which, though hot and limpid, are characterised by high density, 
^uch slags are not conducive to good settling and separation of mattes, 
.and they tend to occasion high copper losses, because — 

{a) The difference in density of slag and matte is not sufficiently 

{b) The solubility of sulphides in the slag increases with its basicity. 

The greater the concentration effected by the smelting operation, 
the higher is the grade of the matte produced ; at the same time, the 
ictual weight of matte is smaller. On the other hand, since more iron 
- oxidised from the charge and slagged off, the quantity of slag pro- 
duced increases proportionately. Contrasting then, the likely losses 
f copper which would result from the association of a small quantity 
"f high-grade matte with much slag, compared with those resulting 
I from the association of a considerable quantity of low-grade matte 
in the presence of but little slag, the former condition is obviously 
the more productive of heavy loss, for not only will many more shots 
of matte be held in suspension, but each shot of high-grade matte 
represents a larger quantity of copper. 

It is found in practice that it is most economical to make a fairly 

' low-grade matte on the first or " green-ore " smelting, and to re-con- 

( <Jentrate this matte pyritically up to converter grade by a second 

^ smelting operation. The extra cost of casting the low-grade matte, of 

Vf^aking up, rehandling it, and resmelting, with all the extra charges 

1' on capital, etc., involved, is less than the losses which would be incurred 
' if higher-grade converter matte were made at the first smelting, 
-although there is no difficulty at all in producing such mattes so far 


as the actual furnace operations are concerned. It is entirely a question' 
of the slag losses involved. 

Under ordinary smelting conditions (not truly pyritic), when using 
some coke for fuel, it would be readily possible to alter the density 
of the slag by adding suitable constituents, such as limestone or 
additional silica, but in pyritic smelting this is not practicable. The 
furnace chooses to make at the tuyere zone its own slag, and that 
a highly basic one. High concentration and a slag low in iron content 
cannot be obtained together in true pyritic smelting, since high con- 
centration means rapid oxidation of iron sulphide, and this necessitates 
high temperature and produces a highly ferruginous slag in con- 
sequence. Additional silica added to the charge could not alter the 
slag composition markedly and still yield the same grade of matte. 
The silica content of the slag depends on the temperature at the 
tuyere zone, and this is governed by the rate of oxidation of the iron 
sulphide. If the slag is to be more siliceous it must be produced at a 
lower temperature, which would be obtained by oxidising the iron 
less rapidly. This would lead to the production of low-grade matte, 
and probably would so reduce the furnace activity that there would 
not be sufficient heat to keep the slag molten. 

If extra silica be added to the charge, it would probably be un- 
attacked unless more iron were oxidised in order to flux it off. In such a 
case the blast would have to be increased in order to produce iron 
oxide more rapidly, the temperature would in consequence be raised, 
a still more basic slag would be produced in larger quantity, whilst 
the matte would be increased in grade and reduced proportionally in 

The addition of sufficient lime to the charge, in order to produce 
a sufficiently low-gravity slag, is also impracticable in true pyritic 
work, because — 

(a) The extra lime consumes silica, and interferes with the desired 
reactions at the bessemerising zone, tending to lower the concentra- 
tion. It also absorbs heat. 

Lime has a very powerful affinity for silica, more strongly marked 
than that of iron oxide, its replacing value is higher, its more siliceous; 
silicates are readily formed and they have a lower formation tempera- 
ture, all of which factors tend to an undue consumption of siUca 
which is urgently required by the iron if the rate of oxidation is, 
to be maintained. The marked tendency for lime in the charge to 
consume the silica tends to retard the oxidation of the iron sulphide, 
which proceeds most satisfactorily when free silica is available for the 
nascent iron-oxide, and in consequence concentration is decreased and 


the heating effect in the furnace reduced. In addition, the larger bulk 
of calcareous slag carries considerable heat from the smelting zone of 
the furnace. Lime silicates and the poly basic lime slags have a 
markedly lower formation temperature than the normal ferruginous 
-lags of true pj^itic smelting, they are hence formed readily without 
lequiring so much oxidation activity at the tuyere zone. In con- 
sequence less iron is oxidised, and the resulting concentration in the 
natte is proportionately reduced. 

ih) The lime is introduced in the form of limestone, and the carbon 
dioxide liberated from this material in the furnace is found to have 
a deleterious effect on the furnace gases if the manufacture of sul- 
phuric acid from them is intended — this being a consideration of 
great economic importance in connection with many modern pyritic 

Hence, in practice, pyritic smelting is at present generally con- 
ducted in two stages for the production of a matte of 30, 40, or 50 per 
cent, converter grade. The " green ore-matte,'' or first matte, runs 
usually from 8 to 13 or 14 per cent, of copper, depending upon the copper 
ore available, which is usually very low grade — 2 to 3 per cent, copper 
contents ; the second or concentrated matte assays 28 to 40 per cent, 
copper. Special care is taken to ensure good settling of matte from 
the basic and irony slags, and by these means the copper losses in the 
-lags are reduced to the comparatively moderate proportions associated 
>vith normal practice. 

It does not appear improbable that with the developments of 
basic converter practice, involving eventually the continuous converting 
of low-grade mattes, the necessity for this second pyritic smelting and 
re-concentration may be avoided. The removal of this feature from 
pvritic smelting practice would add enormously to the potential 
economies arising from the method. 

In spite of the difficulties connected with the process, as detailed 
I hove, the method has proved itself an exceedingly profitable one 
on a large scale, and the experience of the companies financially 
interested, as well as the opinions of managers of the plants in practical 
'peration, leave no doubt as to the economic success of this application 
"f scientific principles to a practical problem on a very extended scale. 

Special Features of Pyritic Smelting. — Several points of par- 

>cular interest have given rise to much discussion in connection with 
J^ritic smelting practice. These include the question of the coke 
roportion required on the charge, and the advisability or otherwise of 
' raploying heated blast for the furnace. 

^'oke Proportion. — Whilst ideal pyritic practice involves the entire 


absence of supplementary carbonaceous fuel, it has not been found 
practicable, up to the present, to ensure satisfactory working over any 
reasonable period of time, unless a minimum of about 1-25 per cent, 
of coke is incorporated with the charge. The function of this coke 
has been a matter of much speculation, but the investigations of 
Sticht already referred to, now permit the tracing, with some con- 
siderable accuracy, of its function and of its action in the furnace. 

It is found that in true pyritic smelting the coke does not reach 
the bessemerising zone at all, but that it is completely consumed in the- 
regions above this point. It is, moreover, not burned by the oxygeu 
of the air, none of which exists above the tuyere zone, since all thi» 
oxygen is consumed by the combustion of the sulphide. It appears- 
that the coke is oxidised by the SOg which results from this sulphide 
combustion. The examination and analysis of samples of the gases, 
withdrawn from different parts of the furnace have confirmed this- 
view, and have elucidated the probable reason for the apparent necessitj^ 
of a certain small proportion of coke in the process, under the present 
conditions of working. The heat generated from the oxidation of the 
coke by the SO^ is of much value in preheating the materials of the 
charge for the removal of excess sulphur and the liquation of the 
sulphides. The amount of heat which is available for this operation 
is small, being practically all derived from that carried upwards by 
the hot gases leaving the smelting zone, and none is obtainable by 
the usual processes of coke or sulphide oxidation in the upper regions- 
of the furnace, since no available oxygen is believed to get past the 
bessemerising zone and reach these upper areas. It is indeed necessary 
for the success of pyritic smelting that such oxidation or roasting of 
sulphides in the upper part of the furnace should be prevented, 
since every available particle of iron sulphide is required for heat 
production at the smelting zone, by its combustion there, and any 
oxidation elsewhere not only deprives this zone of fuel, but spreads 
the heat over too wide an area for sufficiently intense combustion. 

Thus, by supplying an additional amount of heat to the upper 
parts of the furnace, where heat is needed to assist in the preparation 
and liquation of the sulphides, the extra coke, in being oxidised by the 
SO2 without robbing the tuyere zone of fuel or air, just fulfils its useful 
purpose at the required place, in such a way as to keep the smelting 
operation running smoothly. 

The presence of more coke than is absolutely necessary for the' 
fulfilment of this purpose is, in addition to its extra cost, of noi 
advantage, and in true pyritic smelting none should reach the tuyere ^ 
zone, since it introduces a reducing influence where the most marked 


oxidising effect is required. By consuming oxygen for its combustion, 
it deprives the iron sulphide of this material, less iron is, therefore, 
oxidised, and the matte is consequently increased in quantity and 
lowered in grade, whilst the amount of iron carried into the slag is ' 

1-25 per cent, of coke is about the minimum quantity with which 
it is found practicable to maintain satisfactory working of the furnace 
under present conditions, 0-5 per cent, has been worked with occa- 
sionally, and none at all over certain short periods of time. The 
average quantity employed is from 2 to 3 per cent., and when about 
5 per cent, is used, coke reaches the tuyere zone and the process ceases 
to be truly pyritic — the reactions and smelting conditions become 
entirely changed. 

Tt does not seem unlikely that, as knowledge of these conditions 
increases and as the mechanism of the process becomes more generally 
understood, modifications in furnace design and blast conditions may 
lead to the successful operating of the pyritic process entirely indepen- 
dent of the use of coke fuel. 

Heating of the Blast. — For true pyritic smelting it has been shown 
in practice that the use of heated blast possesses no advantages ; many 
smelters operating the process have tested the effects, and have usually 
given the method up, whilst the work of Sticht and Peters affords 
valuable evidence and close argument as to the reasons for its unsuit- 
ability. Success in true pyritic working depends upon the intensity 
of oxidation of the sulphides, and upon the localisation of the resulting 
heat at the narrow bessemerising zone situated just above the tuyeres. 
The greater the quantity of iron which is there oxidised per minute, 
the better is the concentration, the greater is the smelting and fluxing 
intensity and the higher is the resulting temperature. Since the 
character and composition of the slag vary in accordance with these 
conditions, depending largely upon the temperature in the tuyere zone, 
the furnace works most rapidly and satisfactorily when slags of high 
formation temperature are being produced. These can only be formed 
if much iron is being oxidised, because iron is the chief fuel in the 
process. The addition of extra heat by warming the blast appears to 
allow of the formation of silicate slags possessing a lower formation 
temperature, such slags are less basic, and consequently less iron need 
be oxidised and slagged off per minute in order to produce them. Less 
jiron sulphide fuel is, therefore, burned, and the reaction intensity at 
le tuyere zone is reduced, so that the necessary heat margin for 
itisfactory smelting may not be attained. The extra heat carried in 
>y the warmed blast may not be sufficient to compensate for that 


which is lost owing to this decrease in oxidation intensity ; the furnace 
consequently tends to work cold, whilst the excess air supply leads 
to the production of over-fire, by the oxidising of sulphides higher 
up in the charge. 

These features are specially interestinof, as they afford one of the 
most marked distinctions between true and partial pyritic smelting. 
In the latter process, the fuel value in the adjustable supply of coke 
at the tuyeres allows of the ready production of any extra heat which 
might be required. The slag composition is, in consequence, more 
independent of the furnace conditions, since the heat required for the 
smelting operation does not depend so much on the formation of slag 
of any particular composition. Sufficient heat is always obtainable by 
coke additions when smelting for any special slag which may be 
desired. Neither is localisation of the heat at the narrow tuyere zone 
so essential in partial pyritic smelting. Warm blast produces a greater 
combustion intensity when employed in oxidising carbon, so that it 
may present advantages, both economic and operative, in partial pyritic 
work, whereas it is distinctly disadvantageous in the true pyritic 

Pyritic Smelting Practice in Tennessee, — The pyritic process 
is operated in Tennessee at two smelters ; that at Copperhill under 
the Tennessee Copper Company, and at Isabella by the Ducktown 
Sulphur, Copper and Iron Co. The ore averages from 2 to about 2| per 
cent, copper, 31 to 37 per cent, iron, 20 to 30 per cent, sulphur, 10 to 
25 per cent, silica, the remainder being earths, including lime about 
6 per cent., magnesia 2 per cent., zinc 2 per cent., and alumina — i.e., 
a heavy sulphide ore with but little excess of free silica available for 
the fluxing of iron. 

Copperhill. — The process is conducted very much according to the 
principles just considered. The Copperhill plant operates seven 
furnaces of the ordinary rectangular water- jacketed type — the general 
features of furnace design being little different at present, whether 
true or partial pyritic practice be conducted. Several important 
devices in detail have been introduced with successful results, and 
the management is distinguished for its pioneer work and experi- 
mental enterprise in connection with the process. The furnaces were 
formerly all 56 inches wide; three of them are 180 inches long, the 
other four being 270 inches. The height of charge is from 10 to 12 feet, 
the capacity of the smaller furnaces 375 to 400 tons of charge daily, 
and a blast of 19,000 cubic feet of air per minute at 50 ozs. pressure 
is supplied to each. The larger furnaces have a capacity of 500 to 
600 tons daily. Many trials have been made to determine the best 



shape for the water- jacketed sections, both broad and narrow panels 
having been employed. In one of the furnaces, curved end-jackets 
were tried, with the object of lessening the production of crusts which 
tend to form at the corners, owing to coldness and reduced furnace 
activity at these points. The advantages expected have not been 
realised, the tendency to crusting has not been lessened, and although 
barring has been rendered easier, the disadvantages of rounded 
<?orner- jackets and their greatly increased cost of construction out- 
Aveigh their advantages, and their use has now been given up. 

An important modification in the form of the tuyeres has been in- 
troduced with the object of furnishing more effectively the necessary 





1 "^ 





Side \/ie^ Bacl< View. 

Fig. 61.— Slotted Tuyeres, 12 inches by 4 -inches (T. C. C). 

large volume of air at suitable pressure, and of increasing the effi- 
ciency at the tuyere zone. Instead of supplying the air to the furnace 
at a number of separated points, it was felt that the closer these could 
be brought together the better. A narrow slot all round the furnace 
for air admission has been held to be the most perfect method, but 
hitherto it has been thought impracticable, though a recent form of 
furnace (not at this plant) has been devised on this system. The im- 
)rovement here has been the use of slotted tuyeres, 12 inches long by 
"4 inches wide, each of which replaces two of the older tuyeres of 3 J 
inches diameter. These have proved very successful, the furnace thus 

IM»equipped handling a much larger tonnage, and it has been decided 


Charging is by side-dumping V-shaped cars, and great care is 
ta^ken in the handHng and distribution of the charges. The furnaces^ 
are fitted with tops of special design, and with elaborate dust- 
catching devices which have been the subject of long and numerous 
experiments ; the special purpose being to allow the taking off of the 
gases below the feed-floor, and to reduce the height of the super- 
structure to the smallest possible proportions, so as to prevent 
excessive dilution (by air) of the furnace gases, which are used for 
sulphuric acid manufacture. The furnace tops were originally of the 
standard form — brick walls supported by steel frame-work. It was, 
however, necessary to damper down the flues in order to obtain 
sufficient pressure to force the gases through the Glover towers, and 
the heat has caused the steel work to warp badly. A low top was- 
tried, using a brick-hned flue at the end for taking off the gases below 
the feed-floor. This was found to be good for charge-dumping and 
general convenience, but it allowed the escape of too much smoke and 
flames, which greatly interfered with the furnace manipulation. In 
consequence the tubular top was used, gradually raised until a suitable 
height was reached. This form has been described on p. 140. jl 

The present practice at Copperhill is to smelt the ore pyritically for*" 
a 9 to 10 per cent, matte, passing the products through the 16-foot 
settlers which are now lined with siliceous copper ore, then tapping 
the matte into ladles which emptj" it into beds of flue-dust. Alternate 
layers of matte and dust are thus incorporated, and yield a porous 
material convenient for the concentrating pyritic smelt which follows. 
This re-concentration is now conducted in a furnace narrowed ta 
44 inches, which has been found specially well suited for the work ; 
the furnace runs fast, smelting sometimes over 800 tons of charge 
per day. The system of working is that of hard driving so long as the 
furnace smelts rapidly. As soon as it slows down, the furnace is 
tapped out and started afresh. The re-concentrating charge contains 
some hmestone in order to reduce the copper losses in the slag, the 
saving effected by this feature being equivalent to 2 lbs. of copper per 
ton of ore smelted. The resulting matte is bessemerised. 

The furnace gases are utilised for sulphuric acid manufacture, 
the acid plant being the largest in the world, with an ultimate capacity 
of 400 tons per day. 

Ducktown. — It was at the Ducktown Company's smelter that the 
first work on pyritic smelting in the district was carried out, and the 
successful development of the process generally, owes much to Free- 
land's early pioneer work, the remarkable results of which led Parke 
Charming to adopt the process at the Copperhill plant. 










5S '>*OOOC<IOOO©(NO'«* 

1— t 






S ^ • • • ^- • • • • 


Number of 


2, 2, '2, 2 



Mr-* : : : : c- : 


Number of 

2, 2, 2, 2, 2, 2, 2 

2, 2, 2, 2 

2, 2, 2, 2, 2 

2, 2, 2, 2, 2 

2, 2, 2, 2, 2 
2, 2, 2, 2, 2 
2, 2, 2, 2, 2 


2, 2, 2, 2, 2, 2 

2, 2, 2, 2, 2 








r-i ^ 


3 ^ ' ' • 3^ ' ■ ' ■ 




i^ ::§:::: : 



§2 • -N • • • • 



ai ■* . 

S" -5 

.Sfcf(M ;(M ;(M ;(M ;(n ; 'csi 




ill ::::::! 













Coko, .... 
Ore A., . . . . 
Ore B., . . . . 
OreC; .... 


Limo rock, .... 
Gi-een ore (low grade) matte. 
Flue-dust, .... 
Quartz (for flux), 


i/oMr« 0/ Charging 

7-8, . 

8-9, . 

9-10, . 
10-11, . 
11-12, . 
12-1, . 

1-2, . 

2-3, . 

3-4, . 

4-5, . 

5-0, . 


The Isabella smelter comprises two furnaces of moderate size, 
17 feet by 3 feet 4 inches at the tuyeres, having a joint capacity of 500 
to 600 tons daily. The furnaces are about 9 feet high, and are \\-ater- 
cooled. Air at onty 20 to 30 ozs. pressure is supplied through 3-inch 
tuyeres. The smelting scheme is somewhat analogous to that adopted 
-at Copperhill, the first smelting producing a 20 per cent, copper matte 
from the 2 per cent, ore, whilst the re-concentration results in a 
converter-grade matte assaying 50 per cent. The coke proportions 
are somewhat similar to those used at Copperhill, being 5-0 per cent, 
for the first smelting, and 3-5 per cent, for the second. The furnace 
management at this small plant is exceedingly efficient, and the 
campaigns are long, it being claimed that the furnace operations 
have never had to be completely stopped on account of crusting or 
gobbing. This is held to be due to the results of special care in feeding 
and charge distribution, the ingenious Freeland charger already 
described being used. The charge is kept low (6 to 8 feet above the 
tuyeres), and is evenly red hot all through. The slags assay 35 to 36 
per cent, silica, 38-8 per cent, iron, and 80 per cent, lime — with 
moderate copper losses. The annual output is equivalent to about 
3,000 tons of metallic copper. An acid-making plant is also attached 
to these works. 

The Manufacture of Sulphuric Acid from Pyritic Furnace 
Gases. — Modern legislative requirements make severe demands upon 
the managements of smelter- works where sulphury ores are dealt with, 
by reason of the disastrous effects of the sulphurous gases upon the 
conditions of life generally in the vicinity. In other cases, litigation 
by neighbouring farmers and others impose restrictions on the amount 
and character of the gases which the smelters are allowed to emit 
from their furnace stacks. So serious has the problem become that 
several smelters have had to cease operations altogether, others 
have been mulcted in enormous costs by law suits, by claims for com- 
pensation, or by the installation of plant and processes which they 
have been compelled to adopt for dealing with the gases. These 
matters have become subjects of historical importance in the develop- 
ment of smelter practice. 

As has been the case in analogous circumstances elsewhere, when 
interference with the uncontrolled dispersion of then-considered waste 
products has often proved of ultimate benefit and a source of much 
profit to their producers, the enforced treatment of highly sulphurous 
furnace gases has in several instances resulted in considerable gain 
to the copper smelters. 

Among the methods which are at present economically practicable 


for dealing with the smelter gases, those of dilution, and of utilisation 
for acid manufacture are the most important. 

The considerations which decide the best course of treatment 
depend on the numerous economic and local factors which are always 
of such prime importance in connection with industrial undertakings 
demanding large capital outlaj^ The installation of a plant for making 
sulphuric acid from the gases largely depends on — 

(a) The technical factor as to whether the composition of the gases- 
is suitable for the making of acid. 

(b) The economic factor as to whether such acid can be put upon 
the market on a satisfactory basis. 

(a) For the successful operation of acid-making plant, as at present 
developed, it is necessary that the proportions of sulphur dioxide in 
the gases shall not fall below a certain minimum, and further, that the 
gases shall not contain more than certain limiting proportions of other 
interfering constituents, such as, for instance, CO.,. It is for this reason 
that the blast furnace operating the true pyritic process furnishes 
gases of the type most suitable for acid manufacture, since by this 
process the sulphur-dioxide is obtained in the gases in the most 
concentrated and the least contaminated form possible under smelting 
conditions. Even under these circumstances the gases are not in the 
least of an ideal composition for treatment, owing to their dilution 
with nitrogen, etc., and the development of the acid-making plants 
and processes adopted for the successful utiHsation of copper blast- 
furnace gases furnishes a record covering many years of very slow 
and costly experiment, marked by many preliminary failures and 
disappointments. These difficulties have now been overcome, as the 
working of the successful plants attached to both of the Tennessee 
copper smelters affords conclusive proof, and the sulphur which 
formerly cost money to dissipate by roasting, now not only acts as 
fuel, but furnishes a very profitable bye-product. 

The requirements for the gases are chiefly the presence of sufficient 
80., and oxygen, and of as little CO^ as possible — factors which depend 
largely on the proportions of sulphide in the charge. The gas for 
the acid plant must be supplied in regular and continuous amount, 
at a specific temperature, and this calls for special care in the 

I smelting operation, furnace manipulation and blast supply, supple- 
mentary air admission, etc. 
About 3-5 to 4 per cent, of 80.^ in the gases delivered at the 
chambers is the minimum proportion for satisfactory working ; COjj 
should not exceed about 5 per cent., and about 60 per cent, or more of 


(h) In addition to the capital charges involved in the acid-making 
installation and the costs of adapting the furnace plant and operations 
to the process, the problem of putting the acid upon the market on a 
satisfactory economic basis is important, particularly in view of the 
competition from other sources The districts which offer a consuming 
area for the large and regular supply of acid from the smelters are not 
unlimited in number, and are probably readily accessible to other 
sources. In view of the costs of production, the distance of the 
smelter from the market is a serious consideration, since freight charges 
on sulphuric acid are high, involving special regulations with respect to 
the form of car and conditions of traffic, and they may readily exceed 
^11 possible profits resulting from the sale of the product. 

In Tennessee the companies were forced to instal acid plants. 
That at Copperhill is the largest in the world ; commenced in 1906, acid 
manufacture began about two years later, after much experimenting, 
and further units have gradually been added. The plant now 
includes two Glover towers, 30 feet across and 50 feet high, 64 cooling 
chambers about 11 feet x 11 feet x 70 feet high, eight cooling cham- 
bers 11 feet X 24 feet x 70 feet high, twelve old chambers 50 feet x 
50 feet X 70 feet, six new chambers 50 feet x 50 feet x 75 feet, eight 
new chambers 23 feet x 50 feet x 80 feet, eight Gay-Lassac towers, 
with complementary tanks, etc. — producing at the rate of 168,000 
tons of 60° B. acid per annum. 

The Ducktown Company's plant was installed in record time, and, 
like the Copperhill plant, comprises elaborate dust chambers and 
flues, with Glover and Gay-Lussac towers of special design and con- 
struction, and enormous acid-making chambers with complex valves 
and fittings. The plant is designed to produce about 160 tons of 
60° B. acid daily. The analysis of the gases suppUed to the towers 
varied during the early working of the plant ; under fairly normal 
conditions the average analysis of the gases delivered is SO., 3-5 per 
cent., COo 3-5 per cent., SO3 trace; the oxygen in the mixture being 
about 80 per cent. The temperature is also apt to vary. Full 
details on these points are not yet available for general service. 

The management of both companies have been successful in 
obtaining particularly satisfactory contracts for the purchase of their 
acid by fertiliser corporations. 



Peters, E. D., "Principles" and " Practice of Copper Smelting." 
Blast-furnace Maniptdation. 

Shelby, Geo. F., "Alumina in Blast-Furnace Slags." Eng. and Min. Joum., 1908. 
Offerhaus, C, "Copper Blast-Furnace Smelting at Anaconda." Eng. and Min. Journ., 

1908, Aug. 7, pp. 243-250. 
Sackett, B. L., "The Granby Smelter Equipment." Aliiies and Minerals, 1910, April, 

p. 524. 
" Operations of the Tennessee Copper Company. " Official Annual Reports of the Oengral 

Walker, A. L., "The Metallurgy of Copper in 1910." Eng. and Min. Joum., 1911, 

Jan. 7, p. 39. 
Austin, L. S., " Review of Metallurgy in 1910." Met. and Chem. Ind., 1911, Jan. 11, 

p. 40. 
Rice, Claude T., "Handling Copper Smelting Gases." Eng. and Min. Joum., 1911, 

Mar. 25, p. 614. 
"Cottrell's Fume Smelter." Min. and Scient. Press, Aug. 26, Sept. 2, 1911. 
Herrick, R. L., "Boston and Montana Co.'s Smelter at Great Falls." Mines and 

Minerals, 1909, Dec, p. 257. 
Harvard, F. T., " Condensation of Fume and Neutralisation of Furnace Gases." Bull. 

Amer. Inst. Min. Eng., No. 44, 1910, Aug. 
"Mineral Industry." Annual. 

Pyritic Smelting. 

Holway, John, "A new Application of Bessemer's Method of Rapid Oxidation, by 

which Sulphides are utilised for Fuel." Joum. Society of Arts, Feb. 1879. 
Rickard, T. A., " Pyrite Smelting." 
Sticht, Robert, "Ueber das Wesens des Pyrites Verfahrens." MetaXlurgie, Nov. 22, 

Dec. 8, 1906. 
Wintle and Alabaster, "Pyritic Smelting." Trans. Inst. Min. and Met., 1906, 

vol. XV., p. 269. 
Nicholls, F. S., "Pyrite Smelting in Tilt Cove, Newfoundland." Eng. and Min. 

Jmim., 1908, Sept. 5, p. 462. 
Wright, L. T., "Pyritic Smelting without Coke." Min. and Scient. Press, 1906, 

Sept. 29. 

■Sulphuric Acid Manufacture. 

Falding, F. J., and Channing, J. P., "Pyrite Smelting and Sulphuric Acid Manufac- 
ture." Eng. and Min. Joum., 1910, Sept. 17, p. 555. 

Freeland, W. H., and Renwick, C. W., " Smeltery Smoke as a Source of Sulphuric 
Acid." Eng. and Min. Joum., 1910, May 28, p. 1116. 



The Bessemerisinc of Copper Mattes. 

Development of the Process— The Converter— Converter 
Lining's— Grade of lYIatte— Operation of the Process 
—Systems of Working-. 

In modern copper smelting practice, matte of '' converter grade/' 
containing from 30 to 50 per cent, of copper, is bessemerised for the 
production of metallic copper. Successful practice depends upon a 
regular and continuous output of matte from the furnace plant being 
available, and upon a capitalisation and resources on a sufficiently 
large scale for continuous operation of the whole of the smelting plant. 
Development of the Process for Bessemerising Copper 
Mattes, — The success of Bessemer's process, which was applied in 
1856 to the production of steel by blowing air through molten cast- 
iron, led to a suggestion for its application to copper mattes and to 
some experiments on the subject by Semenikow, a Russian engineer, 
ten years later. It was not until 1878 that any further work was 
conducted on a practical scale. In that year John Holway suggested 
and worked out the scheme already referred to, the principles of which 
as outlined by him, form the foundation of the pyritic and converter 
practice of the present time. Air was blown through heated Rio 
Tinto pyrites in an ordinary Bessemer steel converter and the experi- 
ments met with considerable success. The apparatus was, however, 
not deemed convenient, as the process worked very intermittently and j 
large quantities of slag were produced which required to be poured off ^ 
at intervals, whilst the position of the tuyeres in this form of' 
converter was found to be unsatisfactory. There are many practical 
difficulties in employing the same kind of apparatus for the con- 
verting of copper mattes as for the bessemerising of cast-iron into 
steel. In the first instance, the final steel product differs but little 
in weight or bulk from the original charge, whilst the process 
produces but little slag, owing to the comparatively small proportions 
of silicon and manganese which require to be oxidised — whereas 
in copper converting, the quantity of slag produced is almost equal 
in weight to the amount of matte originally charged, whilst the 



resulting copper product amounts to less than one-half of this 
weight. Further, in bessemerising cast-iron, the blow is of very 
short duration ; in copper matte converting, it occupies more than two 
hours, and the relative heat losses are, in consequence, markedly 
different. Finally, the lining of the steel converter chiefly serves to 
protect the shell ; its function in the copper converter was to act 
also as flux for the iron oxides produced on blowing. 

In Holway's final form of apparatus for the pyritic smelting of 
copper ore to metal, the introduction of siliceous material as a flux 
for the iron oxide and the use of basic lining were arranged for, with 
the object of overcoming the difficulties caused by the corrosion of 
the siliceous lining which acted afe flux. 

Though several years elapsed before the pyritic treatment of 

ore was successfully conducted, the process of bessemerising the fluid 

matte to metal was successfully applied on a commercial scale by 

I' Manhes in 1880, although it was not until the following year that 

David's device of placing the tuyeres horizontally and at such a height 

I above the bottom as not to interfere with the metal which is obtained, 

■ solved the final difficulties of operation on a practical scale. In 1883-4 

the Manhes converter was introduced into the United States, and at 

about the same time the barrel form was designed by Manhes and 

David, and was also readily adopted. Both forms developed in size, 

increasing in capacity from 1 ton to that of 7 to 10 tons. 

Until comparatively recent years, the chief modifications in practice 
were concerned with operating and constructional details rather than 
with radical changes in the principles of work. Experiments and 
research have meanwhile been in constant progress with the object 
»f overcoming several of the grave defects connected with the 
apparent necessity for the destruction of the siliceous converter-lining 
hy using it as flux, which was due to the difficulties of causing the iron 
•xide to flux with silica when introduced in any other way. 

The most vital improvement introduced into converting practice, 
ind that with which the future developments are most closely bound, 
s the successful adaptation of basic material for the purpose of lining 
the converter. This achievement, together with recent success in the 
introducing of siliceous flux, promises to solve many of the difficulties 
connected with the bessemerising of low-grade naxtte by a con- 

Kuoits process. 
Suggested by Holway, basic linings were tried at the Parrott 
lelter, Butte, in 1890, by Keller and others, but under the conditions 
ot working at that time they were found to be unsuccessful when 
operated on an industrial scale. Valuable pioneer work was undor- 



taken by Baggaley in Montana, and after many trials, his method 
was successfully operated for some months at the Pittsmont Smelter 
under Heywood's direction in 1906. Visits of inspection to this 
smelter in 1908 proved disappointing, it being found that most of the 
plant which had promised the solution of such difficult problems 
had been dismantled, largely owing to economic difficulties connected 
with its operation, and the works were in process of re- organisation for 
the older system of working. Meanwhile, since 1903, Knudsen, at 
Sulijtelma, Norway, has successfully employed a small basic-lined con- 
verting furnace for the combined pyritic smelting and converting of 
heavy sulphide ores. The process consists usually of pyritic liquation 
of the sulphides, followed by a further concentration of the matte 
up to ordinary converter grade by bessemerising, the higher grade 
matte being then transferred to a silica-hned vessel and blown to 
metal in the usual way. 

The successful operating of the basic-lined converter on the large 
scale and under the conditions of working at great modern plants was 
first established by Smith and Pierce at the Baltimore Copper Com- 
pany's Smelter, and the method has since been installed and worked 
with success at Garfield, Utah (five converters in operation, one in 
reserve) ; at Perth Amboy, N.J. ; at the Washoe Smelter at Anaconda — 
where the whole plant is being adapted for basic-converting — and at 
several other works. 

A recent and promising development has been the reported success- 
ful blowing of fine siliceous concentrates through the tuyeres of con- 
verters at the Garfield Smelter, a method by which it might be 
possible to effect the rapid and efficient extraction of values from fine 
material otherwise difficult to deal with, affording at the same time a 
means of conveniently supplying sihceous flux in a manner possessing 
many advantages. 

Principles of the Bessemerising Process. — The principles under- 
lying the converter process are those which form the basis of pyritic 
smelting practice — of which bessemerising is but a phase. The 
reactions involve the very rapid oxidation of iron and sulphur under 
practically ideal conditions, and the fluxing by silica of the iron oxide 
so produced. The heat of oxidation keeps the materials in a 
thoroughly molten state, and maintains the temperature well above ' 
that required for slag formation and perfect fluidity. The heat i 
derived by the combination of oxygen with the iron and sulphur and i 
that of the iron oxide with silica is developed so rapidly and in such i 
quantity, owing to the large masses now worked with, as to cause a 
reaction-activity sufficient to make the process independent of heat 
from external sources. 


It will be noted how markedly the more recent developments of 
copper smelting have taken advantage of the factois of the time 
element and lyiass influence in obtaining enormous heat intensities and 
consequent high temperatures, by conducting oxidation of sulphides as 
rapidly and in as large mass as possible. The same absolute quantities 
of heat per unit weight of charge were hberated in the older smelting 
methods involving roasting, but the more leisurely manner of operating 
allowed the dissipation and dispersion of much of this heat, thus 
necessitating the employment of supplementary carbonaceous fuel. 

The Converter* — The converter is a lined steel vessel in which 
the molten matte is contained, and which allows of air being blown 
through the material by means of tuyeres which pass through the 

The early form of converter was bottom-blown, and similar to that 
invented by Bessemer, but it was not successful in operation on the 
small quantities of copper matte worked with, owing to the chilling 
effect of the cold air on the copper, which, when produced, sank to the 
bottom and set above the tuyeres, stopping the air blast, and causing 
much loss of metal in the slag. 

The later form of converter was barrel shaped, with a horizontal 
^o^^ of tuyeres situated at some distance above the bottom so as to 
allow the copper to settle, protected from the action of the blast, and 
also to allow of the punching of the tuyeres as required. 

The modern forms of converter comprise both the vertical and the 
barrel types, modified largely as regards size and constructional details, 
and although the vertical form is still in use and is even preferred at 
-^'veral smelters, it has been largely superseded at most plants by the 
warrel-shaped variety, whilst the possibilities of greatly enlarged 
essels using basic linings are likely to favour this replacement still 

1. The Upright Bessemer Vessel is used, and found satisfactory at 
Great Falls and at Mt. Lyell. The general size has been 8 feet 
diameter and 16 feet height, with a capacity ranging from 5 to 12 tons, 
according to the condition of the lining, though at Great Falls con- 
verters of 12 feet diameter with corresponding capacity are now in use. 
f The advantages of the vertical form are, that, owing to the greater 
* depth of matte through which the air passes, the oxidation is more 
rapidly conducted, the lining is more efficiently supported, and the 
wear by abrasion upon the lining is found to be considerably less in 
amount and to be more uniformly distributed. 

On the other hand, the greater depth of matte necessitates a greater 
blowing pressure in order to force the air through the material, 



whilst control over the operations becomes a matter of greater 

2. The Barrel Form of Converter is the type in common use. Among 


Fig. 6*2. — Sectional Elevation and Plan of Barrel-Shaped Silica-Lined Converter (Peters)i 




the advantages claimed for this form are those which accrue from 
being able to operate the same weight of matte in more shallow layers, 
a-s compared with the upright form — thus requiring lower blast 
pressures. Another advantage is the greater ease of regulating the 
depth of material blown through, by tilting the converter and thus 
altering the relative position of the tuyeres. 

Owing to the successful adoption of the basic lining, the barrel type 
of converter has now to be divided into two classes, since the basic 
converter differs from the siHca-lined type in constructional details, 
and is usually of much larger dimensions. Its operation is also con- 
ducted on somewhat different lines. 

Fig. 63.— Latest Form of Silica-Lined Barrel Converter. 

(a) The silica-lined barrel converter varies somewhat in size, the 
Anaconda converters were, however, representative of tlie most 
convenient dimensions. 

The shell consists of |-inch boiler plate, 8 feet in diameter, and 
12 feet 6 inches long. The converter is constructed in two portions, 
e body and the hood, in order to facilitate removal, relining, and 
;eneral repairs. The ends are hned with 9 inches of firebrick, and the 
y with 4 inches ; it is then rammed with lining material to a thick- 
of about 18 inches in all parts. There are 16 1-inch tuyerea 
laced horizontally, and in the latest forms of converter, the air is 



supplied by individual tuyeres which are connected to the blast box, 
and which are provided with ball-valves to prevent leakages and 
back-running during the necessary punching. The cavity is about 
8 feet X 4 fe^t by 6 feet deep when first made, and the converter then 
holds conveniently about 7 tons of matte. The weight of lining is 
about 16 tons, and it lasts six to nine blows. The blast-pressure used 
is 16 lbs. per square inch. 

The hood is bolted on to the body, and is furnished with 
conical safety-pieces to give notice of the wearing through of the 
lining. The converters tilt upon rails, which are strapped round the 
body, and which travel upon rollers. Motion is communicated to the 
converter either by connection with an electrical drive, or very often 
by hydraulic power connecting through a rack to a pinion attached 
to one of the trunnions. The air supply is usually from piston-driven 
blowing engines, communicating through a blast pipe to the hollow 

Fig. 64. — Longitudinal Section of Basic-Lined Converter. 

supporting trunnion of the converter, from which the air passes to the 
blast box. 

(b) The Basic-lined Converter. — The adoption of basic linings is of 
such recent date that although the present form appears to have given 
satisfaction, later developments in basic practice may cause further 
modifications in design. R. H. Vail gives the following details : — 

As at present operated, the basic-hned converters are long barrel- 
shaped vessels consisting of a jf-inch steel shell, 23 feet long and 10 feet 
in diameter, lined with magnesite materials so as to leave a cavity 
about 20 feet x 7 feet x 6 feet. Air is supplied from thirty-two IJ-inch 
tuyeres, each separately connected with the blast box and controlled 
by a valve. Provision has to be made for the marked expansion of 
the basic lining-material by leaving the top of the steel shell open, 
joining-up the free ends by tie-rods (13, Fig. 65), whilst the tuyere-pipe 
connections are flexible The main opening or throat, for the charging 



of matte and flux, is situated in the arch at one end of the con- 
verter ; it is 40 inches in diameter, and surmounted by a short chimney- 
cap of iron, which is 30 inches high and Hned inside with clay. The 
vessel is charged through this opening. Metal and slag are poured 
from the converter through an opening in the side opposite the 
tuyeres, which is kept closed by bricks during the operations. An 
oil-burner is provided at one end, for the purpose of supplying such 
extra heat as might be required, in consequence of undue cooling of 
the copper towards the end of the blow or for heating up the lining 
after repairs. The converter is supported as in acid practice, though 
a tilting device employing wire ropes attached to hydraulic plungers is 
noA\ being introduced in place of the rack and pinion method. 

Fig. 65.— Basic-Lined Converter, indicating Tuyeres, Lining, etc. 

Converter Linings. — The question of the lining has been the most 
important consideration in copper matte converting-practice. 
The functions proper of the hning material are — 

( 1 ) To preserve the steel shell and form a permanent receptacle for 
the molten materials ; by reason of its refractory character. 

(2) To prevent undue losses of heat from the materials ; by reason 
of its low conducting power. 

The employment of the hning material as a provider of suitable 
siliceous flux for the iron oxide, though until recently of vital import- 
ance for the practical operation of the bessemerising process, has been 
a necessary evil in many cases, and although it might have been a 
source of considerable profit under certain conditions, this function is 
unhkely in the future to be the consideration of greatest moment. 
I^ft The vital requirements in modern converter practice are per- 
^^■manence of the Hning and efficient means of effecting the fluxing of 
^^■the iron oxide produced in the converting operation. The necessity 
^^Hfor the frequent refining of converters involves not only heavy direct 
'^■expenses, but it occasions waste of heat in the old finings, waste of 


material, loss of time, interruption of the processes, liabilities to out- 
breaks from the converters, and necessitates much heavy machinery 
for the conveying of vessels for relining, as well as large capital outlay 
in relining shops, plant, and appliances. In consequence, the employ- 
ment of siliceous lining material as flux is usually a most expensive 
method of supplying the requisite silica ; and so much is this the case, 
that an arbitrary limit to the iron contents of the matte has been 
rendered necessary, in order to prevent too much of the lining material 
being used up at a single blow. It was found cheaper to use other 
means of concentrating low-grade matte to a suitable grade for 
bessemerising — i.e., to flux off the excess of iron by means of silica 
in the blast- or the reverberatory-furnace processes. 

Siliceous Linings. — Until recently, the only method for fluxing the 
iron in bessemerising, found practicable on a commercial scale, has 
been by the destruction of the siliceous lining, minimising the 
dead losses as much as possible by employing for the purpose 
siliceous materials from which values in the form of gold, silver, or 
copper could be simultaneously extracted and collected in the products 
of the operation. 

Numerous attempts were made to effect combination of the iron 
oxides with silica introduced by some other method, but none met 
with success. Manhes blew sand through the tuyeres, and obtained 
as result a spongy unfused mass in the converter — whilst silica intro- 
duced in the form of lumps rose to the surface unchanged. In each 
case what silica was required for flux, was taken up from the siHceous 
lining. Experiments of a similar nature, in which basic linings were 
worked with, resulted in the fluxing silica being unabsorbed as before, 
whilst the iron which was in process of oxidation, not finding a suitable 
flux, became super-oxidised, resulting in the production of very 
infusible masses of magnetic or ferric oxides which rendered the process 
unworkable. Baggaley and others in Montana devoted much attention 
to experiments on different methods for introducing silica which would 
flux successfully, methods such as superheating or introducing silica 
held in suspension in fused silicates being tried, but without marked 
success, and for mi any years siliceous linings were necessarily worked 

Owing to the large quantities consumed, the siliceous material 
must be obtainable cheaply and in abundant quantities. It should 
be high in free silica contents, since this constituent alone is effective 
as flux ; it should have the property of binding well with clay or 
other material, so as to yield a rigid and impervious lining ; and most 
important of all from the economic standpoint, it should carry values, 


since by this means only, could its destruction become an actual source 
of profit. At first barren quartz and barren clay were largely used 
for linings, but practice gradually developed in the direction of 
employing more profitable materials, and especially those from which 
the extraction of the values might present difficulties, in treatment by 
ordinary smelting methods. The practice as followed until recently at 
Anaconda is typical of such progress. Until 1908 the lining was 
chiefl}^ made from highly siliceous ore obtained from Snowstorm, Idaho, 
carrying 80 to 85 per cent, of SiO.,, 4 per cent, copper, as well as gold 
and silver, and a little iron and sulphur. This ore was crushed in mills 
and mixed with sufiicient slime from the slime ponds of the concen- 
trating plant to make a binding mixture. The slime, which carries 
about 60 per cent, of silica and also 2-5 per cent, of copper has excellent 
binding properties, owing to its clayey consistency. The proportions 
employed were 3 of siliceous rock to 1 of slime — no water was used, 
the mixture being almost dry to the touch. Since May, 1909, instead 
of employing ore obtained from outside sources, siliceous second-class 
Butte ore, which was formerly concentrated, has been very largely 
incorporated in the mixture used as lining material . it contains 
65 per cent, sifica, about 3-5 per cent, copper, a httle gold and silver, 
and also iron and sulphur. The lining mixture consisted of 2-9 parts 
of this material with 1 part of slime. It was thought at first that 
owing to the greater proportion of sulphides and the lower silica 
content of the Butte ore, this lining mixture might prove inefficient 
compared with the former material, but with somewhat greater care 
in fining, it was found that very little more ore was required, and 
that tested by comparative silica contents it was more effective. Thus, 
where the former linings lasted for an average of six 7i-ton charges, 
equal to 20J tons of copper per lining, the new ones last 5\ such 
charges, equivalent to 17f tons of copper per lining, showing that 
although the efficiency per lining was reduced to 90 per cent., yet, 
alculated on comparative gihca content, the new lining proved to 
be the more efficient. 

The operation of lining is conducted with much care : the old 
lining is knocked away where necessary, rods are placed through the 
tnvf re holes, and lining mixture is dumped in ; 6-inch layers of 

I^niaterial at a time being stamped down hard by means of an Ingersoll 
^■urgent tamping machine, until the lining reaches within 6 inches of 
^^e tuyeres. The wooden mould for the cavity, made up of a number 
of jointed pieces, is then placed in position, and the ramming of layer 
l^hter layer round the sides is continued as before. The hood, inverted, 


converter body and bolted down, a joint being made of moistened 
lining material. The whole operation takes about 1| hours. The 
converter is then slowly dried by a wood fire, coal being subsequently 
added and kept burning under the action of a low blast for five 
or six hours ; it is conveyed to the stand when required, dropped into 
position on the trunnion bearings, and the connections and adjust- 
ments very readily made. 

The manipulation of relining at the Tennessee smelter is conducted 
in a very similar manner. 

Basic Linings. — The all-important feature of the basic lining is 
its permanence, which, rendering the frequent relining of the con- 
verter unnecessary, allows of many economies in connection with 
capital outlay on plant and in operating costs. Further, owing to the 
lessened need for lining repairs, the frequent hauling of converters 
to the repair-shops situated at the further end of the buildings is 
avoided. This allows the employment of much larger converter units, 
with obvious attendant advantages, whilst it increases the ultimate 
possibihty of continuous operation. Thus, the size at present employed, 
though the process has been in operation but a short time, is 26 feet 
by 12 feet, with a capacity of 35 to 45 tons of matte, and a daily 
output of 33 tons of copper from 40 per cent, matte. Such a converter, 
lined with 9 inches of basic material, will operate for 2,000 to 3,000 tons 
of copper before requiring repairs. 

Keller's report on basic linings in 1890 stated that they could not 
be employed successfully, because (a) basic material, being a good 
conductor, caused the outside of the converter to become too hot and 
the inside too cold ; (h) such material broke up easily and so was 
unsuitable for use in permanent linings ; and (c) even when basic 
linings were employed, the silica which was added as flux, refused 
to combine with the iron oxides. These views were very generally 
accepted for some years, until Baggaley's persistent efforts and 
finally those of Pierce and Smith showed that by perfecting the 
constructional methods and details, by preventing heat losses as much 
as possible, and by operating on very large masses of hot material, the 
above difficulties could all be overcome and the basic lining success- 
fully employed. The lining is of magnesia brick, and is 9 inches in 
thickness, except at the tuyeres, where the bricks are 18 inches thick. 
In the bottom of the converter and extending to within 18 inches of 
the tuyere level is placed a filling of ordinary firebrick, which is 13| 
inches thick in the middle and 4 inches thick at the sides. The 
magnesite bricks are laid in dry magnesite powder, except near tbei 
tuyeres, where a mixture of magnesia and linseed oil is used. 



Expansion cushions of wood are inserted at intervals along the side 
of the fresh hnings which are then ''seasoned'' with molten copper. 

The required quantity of siliceous flux, as calculated, is now 
successfully introduced by dumping it into the converter, and 
pouring the matte charge upon it. 

The Grade of Matte for Converting.— The grade of matte which is 
economically the most profitable to treat in the converter is a factor 
of great importance, since, if Hmits be fixed, the prehminary 
smelting stages for matte production are made less flexible, whilst in 
order to obtain matte of the correct grade, the smelting oper- 
ations may require to be conducted at greater cost, or else additional 
smeltings for further concentration of the first matte may be 
necessitated — as is the case, for instance, in pyritic smelting at present. 

The grade of a matte is usually expressed in percentages of copper, 
but from the standpoint of the practical converter operations, the 
proportion of iron is the factor which decides the suitability or 
otherwise of the matte for treatment, and since mattes may be 
regarded as mixed sulphides of iron and copper, a matte rich in 
copper is correspondingly low in iron contents, whilst a low-grade 
matte is high in iron. 

The importance of the iron contents of the matte from the view- 
point of converter practice is due to iron being the chief source of 
heat in the operations, and to the fact that the iron oxide produced 
from it is the constituent which requires a supply of flux in order that 
the reactions may proceed and the process be successfully operated. 
The economic limit to the grade of matte suitable for the converter 
process is reached when it becomes less costly and more profitable to 
supply the required siliceous flux for the iron in the ordinary smelting 
furnace rather than in the converter. So long as the destruction of 
the lining was practically the only medium by which silica could be 
efficiently supplied, the limit to the iron contents of the matte was 
fairly rigid. 

The besseme rising of a low-grade matte (low in copper contents, 
high in iron) entails the great advantage that a high temperature 
is obtained, owing to the fuel-value of the iron. On the other hand, 
however, grave disadvantages attend such practice, especially when 
working with the comparatively small quantities of material usually 
perated, and when employing siliceous linings. These disadvantages 
elude the factors that — 

(a) Large quantities of iron oxide are formed, which require siliceous 

ih) Large quantities of highly ferruginous slag are produced which 



carry copper values, and which also demand special attention in 

(c) The quantity of copper obtained is comparatively small, thus 
increasing the proportionate losses and working difficulties. 

In bessemerising a high-grade matte, the heat production is much 
smaller, owing to the decrease in the quantity of iron, which is the 
chief fuel of the process, and the limiting grade is quickly reached 
above which the bessemerising operation upon the matte ceases to be 

In consequence, up to a comparatively recent date, a compromise 
has necessarily been effected, and the grade of matte operated upon 
has been such as to cause as much heat production as possible, together 
with the smallest practicable amount of fluxing action. 

On these grounds, a matte containing from 40 to 50 per cent, 
of copper (equivalent to 32 to 22 per cent, of iron) has been found 
generally the most suitable. At several smelters, lower-grade mattes 
of from 32 to 40 per cent, copper- contents are converted most 
profitably, owing to such special circumstances as the profits resulting 
from the destruction of lining material, or in consequence of the 
fact that greater operating costs would be involved in concen- 
trating the matte to a higher grade by the ordinary- furnace-smelting 

In this connection, the successful adaptation of the basic lining by 
permitting the supplying of flux by means other than from the linings, 
has very important application and possibilities. 

Owing to the frequent relining of these converters being then no 
longer necessary, mechanical difficulties of conveying the converter 
bodies to the relining shops are lessened, and larger converter units 
can now be employed, treating, even at the present stage of develop- 
ment, between six and seven times as large an amount of matte as 
formerly. By operating on such big charges, pouring off slag as 
produced, and adding fresh matte and flux without fear of destroying 
the lining, the difficulties attending the converting of low grade mattes 
have been successfully overcome 

The limit to the grade of matte economically suitable for the 
process will depend, in the future, chiefly upon the comparative costs 
of effecting the required concentration up to any desired grade, in the 
blast- or reverberatory-furnace, or in the converter. 

The modern smelting scheme appears, therefore, likely to develop 
into the preliminary smelting of the ores by the cheapest method 
available, for matte of a grade best suited economically to the 
running of the furnace, the grade being independent of any rigid limit 



for the subsequent converting operations — the matte being then 
bessemerised as usual. 

The Converting Process — Acid Lining.— There are t\\o main stages 
in the converting of copper mattes. The first is essentially elimination 
of iron sulphide ; the second, elimination of the remaining sulphur. 

The product of the first main stage is a white metal, practically 
pure copper sulphide, the iron of the matte having been slagged off 
in the form of silicate, and the corresponding sulphur eliminated as 
SO^. The reactions during this stage are well known : the oxygen 
of the air blown in, yields oxides of iron and of sulphur, as well 
as some copper oxide. The latter, immediately reacting with iron 
sulphide which still remains, re-forms copper sulphide, with the 
production of more iron oxide. The iron oxides are fluxed by the 
liceous materials present, forming ferrous silicate slags. The iron 
uxidation is productive of the greater part of the heat in the operation, 
and high temperature usually marks this stage of the process, which 
may be termed ''the slagging stage.'' 

The flame which issues from the converter during this period is 
usually characterised by a green colour, caused apparently by the 
! formation of iron-sihcate slag. 

When this stage is completed and the slag poured off, the white 
metal is blown up to blister copper — this constituting the second main 
tage of the process. The chief reactions are those of sulphur elimina- 
ion and the production of metallic copper, caused by the action of 
')me of the copper oxide first produced, upon the copper sulphide 
nil present. 

The flame during this period is small, thin, and fairly non-luminous, 
usually of a red-purple to bronze purple colour.* 

The progress of the blowing from copper matte to white metal 
and thence to blister copper is usually indicated and controlled at 
the smelter by the appearance of the flame which issues from the 
ioae of the converter during the first periods, and by the character 
)f emitted shots during the later stages. This is particularly the case 
vith mattes of moderate purity worked in the siUca-lined converter, 
rhe successive changes in these indications are gradual, but are 
•a«ily followed by the experienced skimmer, who is thus able to 
iudge readily as to the manner in which the blow is progressing, 

Kid also as to the temperature, composition, and nature of the metal in 
e converter. 
In general character, this colour sequence, during the bessemerising 
* With nickeliferous mattes, the green colour in the flame in reported by Heywood to 
Riwt throughout the whole process. 



"* -^ -^ O CO 

9000 O 
6 6 6 6 6 

o 00 









CO --• t^ 

O O O O —I 
C5 -^ (M O O 

6 -ri 6 S 6 


So 00 
o ^H CO CO 

s 6 -^6 

2 g 

UO QO 00 

t^ S (M 













•M -^ n Q 

00 ^ t- J5 


r^ 't* -H X 

t- t- X C5 



? § 

i^g§ 2 









U) bO 

























^ ^ 












— 1 












































o -e 

► bo bo bO 

bO fi fl C 

? ^ _0 _0 : 


1 I 


5 ^ ^ o^-C § ^ ^ ^ ^ 


"33 "5 

o -. w 

^ ^ „ _ „ _ _ 

0Ss0"-0 — SS2 

;:;: tep:r?-^rj c3 ce c8 ee 



-C -^ -C 

c g c 

3 S =5 

© s © 

P-i :3 2 

^ © 

03 " 

^ ^ ^ 


.j--g ^ o 

-fcS .,H .3 ♦^ 
Sh -i-< -1^ _« 


ce PHP^pH{iiP^aPHPH&&afi^ 


Ph a a Ph ai 

S S S S £ S S S S S S 


-M 'M iM (M (M iri 


of the ordinary class of copper mattes — i.e., those consisting largely of 
iron, copper, and sulphur, with but moderate quantities of impurity — 
does not vary very markedly, but the body and luminosity of the 
flame depend to a great extent on the nature of the charge and 
on the working conditions. The colours are intensified by very hot 
metal, large charges, heavy blast, and rapid working, and particularly 
by the presence of secondary constituents, such as zinc^leaiJLo^arsenic, 
which liberate dense white fumes, and so increase the luminosity oi 
the flame. 

There are generally four main variations in the appearance of 
the flame from the acid-lined converter : — 

( rv • J <f c J J.-.. J. J l>ail< reddish-brown Haine 

, , , , ) Oxidation of secondary constituents, f . . , , 

At commencement of blow, ■( • .• i u j i > Accompanied bv much 

) Burning of iron, sulphur, and coal, ^ 

\ ) smoke. 

Slagging stage, . . Iron-sulphide oxidation, . . Apple-green flame. 

White metal stage, . . Copper oxidation in presence of slag, White- blue flame. 

Blowing to blister copper, Sulphur oxidation, . . Thin red-purple flame. 

The changes in composition of the charge during a converter blow 

have been traced by Mathewson, who assayed samples during the 

j various stages ; some of these results are indicated in Table xii. and 

I in Fig. 66. For full record see Trans. Amer. Inst. Min. Engineers, 1907. 
In general, of the constituents present in the matte, iron and 
' sulphur are removed very readily, 96 per cent, of the former and 
i 53 per cent, of the latter in the slagging stage of the blow, whilst the 
|: elimination of the injurious impurities is high, bismuth and arsenic 
\ being removed to the extent of upwards of 90 per cent., and of the 
I antimony, selenium, and tellurium, from 40 to 70 per cent, are 
' eliminated (see p. 217). 

Working of a Typical Charge in Silica-lined Converter. — The Ana- 
conda converter plant is now being operated with basic linings. The 
j former practice at this works was representative of the best type of 
\ acid-lined working, and the following description, based upon this 
: practice, is typical of the method in general use. There were in 
operation twelve converter stands of the dimensions previously given. 
Normal working was to convert the 45 per cent, copper matte to white 
metal, to pour off slag, blow to blister copper, and pour the resulting 

|»netal — in regular sequence. 
The colour-changes in the flame during bessemerising are indicated 
the colour-photographs reproduced in the frontispiece. 
•Seven to 8 tons of matte at an average temperature of 900^ C. 
are charged into the converter, which is in an upright position with 
the blast on (16 lbs. per square inch). The operation of charging 



occupies three minutes. A few lumps of coal are thrown in, a vigorous I 
action commences, copious and heavy white fumes and smoke and a 
full red to red-brown flame being emitted. The converter is now 

5 65 





-ji SO 

6 « 

^ 40 



























s, — 



// r\ 


























JO 20 30 40 50 60 70 80 90 100 110 120 130 140 

Time of Blowing in Minutes. 
Fig. 66. — Composition of a Charge during Bessemerising Operation. 

turned slowly back, so as to bring the tuyeres more completely under 
the charge and ensure more rapid and efficient oxidation, and the 
blow proper then commences. The flame drops for a time, continuing 



to be of a red to red-purple colour for two to eight minutes, after which, 
green commences to show in the red smoky flame (A), indicating that 
the first or slag-forming period of the blow is beginning. The green 
colour becomes more prominent and continues for 40 to 45 minutes (B). 
A preliminary pouring off of slag is then usually made, owing partly to 
the danger of violent or even explosive interaction which might other- 
wise occur between matte and slag, and also with the object of keeping 
do^^^l the copper losses in the slag by removing the greater portion of 

Fig. 67. — Pouring Slag, Anaconda. 

f he latter at as early a stage as possible. The blowing is then continued. 

i'lashes of blue now occasionally appear in the flame, and gradually 

:rease in number until the flame becomes blue-white (C), which 

cates that most of the iron has been slagged off and that the white 

I stage is reached. The blue-white colour of the flame is to be 

tributed to the production of copper silicate, owing to the tendency 

the copper oxide formed by the air blast at this stage, to flux off, 

nd to produce the silicate rather than attack the copper sulphide. 

I'his formation of copper-sihcate is particularly liable to occur in the 





presence of much slag and at high temperatures, factors which are 
well known to encourage this selective combination, and which prevail 
at this stage. 

The blowing up to white metal takes about one hour.. 

Slag is then poured off again, until an iron rabble heW under the 
stream commences to show signs of " metal '' which give an appear- 
ance of spots of grease on the blade. The charge is then usually 
** doped.'" '' Dope " consists of highly cupriferous scrap, cleanings, 
slags, residues, also some siliceous material, added partly for the pur- 
pose of cooling down the charge which tends to become overheated 
at this stage. The converter is turned up again and the blowing 
is resumed in order to convert the white metal to blister copper. 

The main reaction which now proceeds is represented by the 
equation ^^^^ ^ ^^^^ ^ ^^^ ^ ^^^^ 

This stage of the blow also occupies about one hour or more, 
according to circumstances. It commences with a vivid red flame 
accompanied by smoke, but this soon dies out and a thin purple, 
almost colourless, flame results, which continues practically un- 
changed for the remainder of the blow (D). The temperature of the 
white metal is to some extent judged by the appearance of the flame, 
a red-brown colour indicating the correct temperature. If the colour 
be too red, the metal is too cool, and coal is thrown in : if the tint be 
too orange, the temperature is too high, and dope is added. Constant 
punching of the tuyeres by long steel chisels is required during this 
stage of the blow, owing to the lessened heat production due to 
diminution of iron, and also to the marked tendency for the 
liberated copper to chill round the tuyeres. The end of the blow is most 
difficult to judge, and although the size and colour of the flame offer 
some criterion, the usual and most important guide is the emission of 
small shots of copper which no longer stick to the hood situated above 
the converter throat, but which rebound from it. This is the stage 
where the skill and judgment of the skimmer are most tried. 

When the blow is considered satisfactory, the character of the 
metal is further tested by pouring a small quantity on to the floor — 
a rugged and uneven surface indicating satisfactory metal. If poured 
too soon, the copper is coarse and impure ; if poured too late, heavy 
losses in the slag result, owing to excessive oxidation of the metal. 

The copper is then poured into a ladle, and conveyed to the refining 
and casting furnaces. 

The whole operation for a straight run occupies about two hours, 
but the time required in general naturally depends upon the rapidity 


of working, and particularly on the grade of matte, and the volume 
and pressure of the blast. 

The slags during the early part of the blow generally carry about 
2 per cent, of copper, after the Avhite metal stage is passed, they 
are usually much richer, on account of the intensely oxidising atmo- 
sphere which prevails, and the decreasing quantity of protecting 
sulphur. These later slags often contain upwards of 20 per cent, of 
copper, and in consequence as much slag as possible is poured off during 
the early stages of the blow, and the quantity towards the close is kept 
at a minimum. 

The subsequent treatment of the converter slag depends very much 
upon the conditions of work at the smelter ; at Anaconda, the iron 
contents of this slag are very useful in the blast-furnace charge, as 
there is a shortage of suitable basic flux for the silica of the rather 
siliceous charges. The slag is poured from the converters into ladles, 
and conveyed to a slag-casting machine, consisting of a conveyor 
belt carrying cast-iron moulds which are sprayed with cold water, the 
slag being thus cast into cakes suitable for the blast-furnace charge. 
At Tennessee, the pyri tic-smelting slags are already too ferruginous 
for any addition of irony converter-slags in the blast-furnace charge to 
be desirable, and the only metallurgical treatment for which these are 
suited is that of recovering from them the large amount of copper 
which they carry. The molten converter slag is, therefore, poured 
directly into the blast-furnace settlers, and by this means, the slags 
re cleaned and the values recovered. 

At the new Tooele Smelter, under Mathewson's organisation, the 
molten converter slags are poured directly into the reverberatory 
furnaces, there being no blast-furnace or settler plant, and the 
^leaning and settling are thus very satisfactorily conducted. 

Systems of Working: Acid-lined Converter.— The "normal" 
vstem of working — i.e., blowing a matte-charge first to white metal, 
lien to blister copper — is not always practicable nor economically 
ue best practice, and the system of operating the charges depends 
iTgely upon the working conditions, which are subject to much 
iriation at different smelters. Even at the same plant, the procedure 
-us to be varied according to the attendant circumstances. 

(Conditions which may influence the system of working include : — 

(a) Grade of matte. 

(b) Temperature of matte. 

(c) Condition of converter lining. 
{d) Rate of production of matte. 

(e) Condition of affairs at the casting and refining furnaces. 


As instances of the way in which some of these circumstances- 
affect procedure, the following examples may be quoted. 

(a) When working with matte of low grade, especially in small 
quantities, as formerly operated, the loss of heat by radiation and by 
that carried away in the large quantity of slag produced, is very con- 
siderable, whilst towards the later part of the blow, the amount of 
sulphide fuel diminishes to such an extent that the maintenance of 
the desired temperature is difficult. The bulk of the final copper- 
product of the operation is very small and the metal is therefore 
liable to chill. In such cases, the system of " doubling '' is useful. 
This consists of blowing the matte to the white-metal stage, pouring 
off the slag and adding a further charge of matte. This, on the 
resumption of blowing, restores heat and yields a charge of white 
metal sufficient to maintain the required temperature for the last 
stage of the blow, as well as affording a convenient yield of metalhc 

(c) i. When working with a freshly-lined converter the charge is 
necessarily rather less than usual, owing to the smaller size of the 
cavity, and this results in a smaller yield of white metal, which is 
also colder. At the white-metal stage the slag is poured off, and the 
cavity having now become larger, owing to the fluxing action upon 
the lining, a fresh charge of hot matte is added, introducing fresh heat, 
further enlarging the cavity, and providing for a hot and plentiful 
supply of white metal for the blowing up to blister copper. 

(c) ii. When the lining commences to wear thin, the converter 
may be retained solely for the purpose of blowing successive charges 
of white metal up to blister-copper, since owing to the very low iron 
content of white metal, there is little fluxing action on the lining during 
this stage, whilst the large quantity of white metal which can be 
operated in the enlarged cavity ensures a good supply or heat. 

When linings bum through, the charge is transferred to another 
converter and the bessemerising finished there. 

The management of the converters as thus indicated, and the 
distribution of the charges among the various converters are left to the 
head skimmer, who has control of the converter floor. 

Working of the Basic-lined Converter. — The actual operations of 
bessemerising in the basic-lined converter differ but little from those 
where the silica hning is used. One important change has, however, 
been made, viz. : the introduction of the sihceous flux before the 
commencement of the blow. The lining having been heated up and 
" seasoned," the charge of four or five ladles-full (30 to 40 tons) 
of matte is poured into the upright converter through the 



throat, 3 to 4 tons of siliceous flux, which must be well dried, are 
added, and the blast is turned on gently (at 5 lbs. pressure), whilst 
the converter is slowly turned back — these precautions being necessary 
in order to prevent excessive blowing out of the dry siliceous fines at 
the commencement of the work. When the sihca is fairly well incor- 
porated, the blast-pressure is increased to about 10 to li> lbs. per 
square inch, the blowing is continued for 30 to 45 minutes, and after 
the silica has been fluxed by the iron oxides — which is tested by feeling 
the charge with an iron rod inserted through an opening in the breast — 
the converter is turned over and the slag poured off. A fresh charge 
of matte and a further quantity of siliceous ore are added and the 
blowing is resumed, these operations being repeated several times 
until the desired quantity of white metal has been accumulated, which 
is then blown up to blister copper in the usual manner. During the 
«arly stages of the blow, the operation is largely controlled by judging 
the quantity of iron remaining in the matte, from the appearance of 
f?mall samples which are ladled oat of the converter from time to time, 
s,nd from this, the quantity of siliceous material required for the further 
fluxing is deduced. This material must be quite dry, so as to flux, 
evenly and not form floaters. One of the advantages of the basic 
process is that siliceous ores containing values (the extraction of which 
may be profitable) which might not be suitable for use in sificeous 
linings, can be conveniently employed as flux in conjunction with 
the basic lining, though naturally the best work is done with flux 
containing a maximum of free silica. The character of the slag is not 
very different to that produced in the siHca-hned converter, though 
it is usually lower in silica contents, and owing to the methods of 
frequent pouring, it is lower in copper values. 

Special Features of Basic-lined Converter Work. — The basic-lined 
converter tends to lose heat by radiation and conduction more quickly 
than does the sihca-lined vessel, due to the walls being thinner and 
the lining material a better conductor. Owing, however, to the use of 
larger charges, to the increased fuel value of the low-grade mattes, 
and to the larger blast-volume used, heat is retained sufficently well 
for the successful operation of the bessemerising process. The tem- 
perature is, however, generally lower than that obtained when using 

he siliceous lining, and constant punching of the tuyeres is necessary — 

o men being required per shift for this work. The great advantages 

of the basic lining are connected chiefly with the fact that the frequent 

refining associated with the silica-lined converter is avoided, hence an 
extensive refining plant is not required, smaller building space and a 

lighter crane can be used. The use of basic finings further affords a 



means of extracting the copper and other values from sihceous ores 
which can be used as flux, but which might otherwise be difficult 
to treat, and it has made possible the cheaper treatment of low-grade 

The disadvantages are chiefly those caused by 

(a) The use of a material which is not perfectly suited for construc- 
tional work, hence repairs occupy longer time. 

(h) The risks of destroying the lining mechanically near the tuyeres, 
owing to the extra punching required at these points. 

(c) The operation and manipulations requiring extreme care and 
attention, owing to the tendency for the production of very high 
temperature during the great evolution of heat in the early stages of 
the blow, when large quantities of iron are being oxidised. 

id) The tendency to losses, by the blowing out of the dry siliceous 
ore, when first turning on the blast. 

Converter Shop Organisation, — The introduction of the basic 
lining has, to a large extent, overcome the necessity for devoting so 
much shop space to the repair department, which formerly occupied 
a very considerable area. The converter stands are usually placed 
in alignment down one side of the building, the centre space is kept 
clear, and is commanded by the travelling crane for the conveyance 
of the ladles of matte, metal, or slag, to or from the converters. At 
Anaconda, the converters are charged from a train of matte-ladles 
mounted on bogies which run along a track behind the converters 
and situated some distance above them, the matte being poured 
down a launder which swings into position over the converter throat. 

At Copperhill, Tenn., the converters are charged from ladles which 
are filled from the blast-furnace settlers situated at the other side of 
the furnace-building, whilst at the most modern large plant, at Tooele. 
Utah, the matte is run directly from the reverberatory furnaces to the 
converters along launders which are nearly 80 feet long and inclined 
at about 7 in 100. This method avoids all the handling of matte 
by cranes and ladles with the attendant troubles of skulls, breaks- 
down, spills, etc., and no difficulty has been found in keeping the 
channel free and open, nor in supplying matte at a sufficiently high 
temperature. At Anaconda and Tooele, the side of the converter- 
shop situated opposite to the converters is devoted to the refining 
and casting furnaces and to the slag-casting machines. 

Modifications of Converter Practice. — (1) David's Best Selectincf 
Process. — David devised a special form of converter and suggested a 
method for conducting in the converter, instead of in the reverbera- 
tory furnace, the operations of the best " selecting process '' on the prin- 



ciples of the old Welsh practice. The method embodied the converting 
of the matte somewhat beyond the white metal stage, by which means 
a small quantity of metallic copper was produced, in which the whole 
of the gold and silver values and most of the impurities collected, 
the remaining white metal being left tolerably pure. The metallic 
copper, thus obtained, was run into a side pocket in the lining and tapped 
from there, the rest of the pure white- metal was blown up to pure 
best-select copper. 

The method is, however, too specialised for ordinary commercial 
copper smelting, especially when electrolytic refining of the crude metal 
can be conveniently arranged for. 

(2) The Haas Converter. — The Haas converter is spherical in form, 
and the tuyere holes through the lining are arranged at such an angle 
as to lessen the pressure required for the forcing of air through the 
metal. It is claimed for this form that it ensures better mixing of 
the materials and more even wear on the lining, by imparting a 
swirling motion to the bath. 


Douglas, James, "Treatment of Copper Matte in the Bessemer Converter." Trans. Inst. 

Min. and Met., 1899, vol. viii., p. 1. 
Baggaley, " A Brief Description of the Baggaley Process." 
Heywood, W. A., "The Baggaley Pyritic Conversion Process." Eng. and Min. Jaum., 

1906, Mar. 24, p. 576. 
Knudsen, E., " Pyrite Smelting by the Knudsen Method in Norway." Mineral Industry, 

vol. xviii. 
Moore, Redick R., "Copper Converters with Basic Linings." Eng. and Min. Jovrn., 1910, 

June 25, p. 1317. 
Editorial, "Improvements in Copper Smelting." Eng. and Min. Jmm., 1911, Mar. 4, 

p. 450. 
Schreyer, Fr., " The Question of the Basic Bessemerising of Copper Mattes." Metallurgie, 

1909, vol. v., No. 6, p. 190. 
" Improvements at the Washoe Smelter." Mines and Minerals, 1910, April, vol. xxx.. 

No. 9, p. 520. 
Vail, R. H., "The Pierce and Smith Converter." Eng. and Min. Jaimt., 1910, Mar. 12., 

p. 563. 
Moore, Redick R., " Basic-lined Converter for Leady Copper Mattes." En/j. and Min. 
Jaurn., 1910, Aug. 6, p. 263. 
"Recent Practice in Copper Matte Converting." Eng. and Min. Joum., 1910, 
Sept. 3, p. 460. 
Neal, Carr B., " Further Data on the Basic Converter." Eng. and Min. Joum., 1911, June 13, 

p. 964. 
KeUer, E., " A Study of the Elimination of Impurities from Copper Mattes in the Rever- 

beratory and the Converter." Mineral Industry, 1900, vol. ix., p. 240. 
Mathews<jn, E. P., " The Relative Elimination of Iron, Sulphur, and Arsenic in BoBSomerUing 
Copper Mattes." Bull. Amer. Inst. Min. Eng., 1907, Jan. 7, No. 13, p. 7. 



Offerhaus, C, " Operation of an Anaconda Converter." Eng. and Mm.Joum., 1908, Oct. 17, 

p. 747. 
Levy, D. M., " The Successive Stages in Bessemerising Copper Matte as indicated by the 

Converter Flame." Trans. Inst. Min. and Met., 1910., vol. xx., p. 117. 
Hixon, H., " Notes on Lead and Copper Converting." 
Sample, Clarence C, " Analyses of Converter Fume." Eng. and Min. Joiirn., 1911, Mar. 11, 

p. 508. 
Haas, Herbert, " The Vortex Copper Converter." Eng. and Min. Joum., 1910, May 7, 

p. 972. 


The Purification and Refining of Crude Copper. 

Preliminary Refiningr and Casting: into Anodes- Elec- 
trolytic Refining" -Bringringr to Pitch, and Casting of 
lYIerchant Copper. 

The further treatment of the converter metal depends to a large 
extent upon its composition, and the purpose for which it is intended. 
The matte-smelting operations on copper ores bring about the 
elimination of the greater part of the constituents accompanying the 
copper. The converter-grade matte may, however, in addition to 
the copper, iron and sulphur, also contain considerable proportions of 
easily reducible impurities of the ore, possessing a greater tendency to 
enter the matte than to be oxidised and eliminated in the slag. Such 
constituents may include gold and silver (practically all concentrated 
and retained in the cupriferous product), arsenic, antimony, bismuth, 
selenium and tellurium (retained to very considerable extent), as well 
as lead, zinc, nickel and cobalt (in much smaller proportions). The 
amount of these latter impurities ultimately retained in the converter 
matte depends very largely upon the proportions originally present in 
the ore, and upon the smelting conditions. 

Under the strongly oxidising conditions of the Bessemer process 
the copper retains but small quantities of impurity, and those which 
do remain in ordinary converter metal may be broadly divided into 
two classes * — (a) those which are oxidisabJe with comparative ease, 
and (b) those which persist in the metal even under oxidising 
influences, unless treated by special means. The former include iron, 
sulphur, and zinc; the latter, arsenic, antimony, bismuth, selenium, 
tellurium, gold, and silver. Keller gives the following figures for the 
average elimination of the impurities in the converter : — 

Arsenic,. 81 per cent. ^ 
Antimony, 71 ,, 
Selenium, 47 ,, 

Iron, . 99 per cent. 
Sulphur, 99 
Zinc, . 99 ,, 
Cobalt, 99 
Bismuth, 97 „ 
Lead, . 96 

Tellurium, 40 
Nickel, . 37 

* The case of nickel may be here treated as exceptional. It is eliminated with difficulty, 
the nickel and copper tending to oxidise together on bessemerising. 



Of the persistent elements, the retaining of the gold and silver 
in the converter-copper is a factor of much economic advantage, 
but the other impurities are curiously just those which are 
characterised by possessing most injurious effects on copper intended 
for electrical work — for which purpose most of the material is^ 

The demand for particularly pure metal in electrical and con- 
ductivity work therefore usually necessitates a further purification of 
the converter-copper (unless it be an exceptionally pure brand) and 
the production of metal specially free from the injurious constituents 
which persist to a small but sometimes very appreciable extent in the 
metal under the ordinary oxidising conditions. The presence of silver 
and gold in the copper may afford in many instances sufficiently good 
reason for a separating process independently of the market for the 
pure copper itself. 

In modern practice, electrolytic methods are almost universally 
employed for the purification of the crude copper. By this means 
the large demands of the present day can be conveniently met, and the 
copper be obtained in a condition of remarkable purity. The frequent 
presence of gold and silver in the metal, and the convenience and 
completeness with which they are separated on electrolytic treatment 
of the copper are particularly advantageous features which recommend 
the adoption of electro-refining, and may in some cases be the reason 
for this procedure even though the metal might otherwise be already 
quite up to specification for electrical service. In the large majority 
of cases, these bullion- values constitute a welcome and independent 
bye-product, the returns from which may be set against the expenses 
of the refining operations on the copper, which might, in any case, 
be necessary. 

The process may, therefore, be operated with one of the following 
objects — 

(a) Of purifying converter-copper. 

(b) Of recovering from copper, the bullion- values which have been 
collected in the metal. 

(c) Of manufacturing pure copper, and recovering the gold and 
silver as profitable bye-products. 

Under the present industrial conditions, the electrolytic refineries 
are located at centres often at very considerable distance from the 
smelters. Situations for the refineries are chosen where the local con- 
ditions as regards power supply, technical resources, and particularly 
proximity to markets and distributing centres, allow of the operations- 
being conducted under the most advantageous circumstances, and it i& 



customary for smelters situated in the remoter mining districts to- 
ship the crude copper to these custom refineries, instead of conducting^ 
the process themselves. At Anaconda, the well-equipped electrolytic 
refineries have been closed down, and the anode metal shipped to the 
Eastern refineries for treatment. 

Preliminary Refining of Converter Copper and Casting into 
Anodes. — For modern electro- refining practice, the crude metal must 
be prepared into anodes, which are usually in the form of plates about 
2 feet 6 inches x 3 feet by 2 inches thick. It is found that the 
metal as produced in the converters, on being cast into such plates, 
does not as a rule yield anodes which work satisfactorily in the tanks. 
This is largely owing to the impure and crude condition of the metal, 
which results in the production of plates which are spongy, coarse, 
and exceedingly rough and uneven on the surface. In consequence 
the direct employment of such metal would occasion irregularity and 
difficulties in the operation of the tanks, giving rise to short circuits, 
uneven wear, breaking off in large pieces, and similar troubles. 
Furthermore, the tank liquors and slimes become badly contaminated 
if large quantities of impurity be present in the anodes, and the 
deposition of good clean metal is thus greatly interfered with. All 
these reasons render it advisable that the converter-copper should, 
as ^a^jui^ undergo a-praLimiiiary furnace treatment before being cast 

The Anaconda practice is representative of the manner in whicli 
these preliminary refining and casting operations are conducted, 
except that the enormous scale and organisation of the operations 
are practically unique. The principles involved and the general 
method of operation are in all essentials those • of the old Welsh 
furnace-refining process. 

The Furnace. — The finished metal from the converters is teemed 
into ladles, and from these is poured directly into one of three 
casting furnaces. Two of the furnaces are in constant use, one of them 
engaged in refining, one being filled, and one in reserve or repair. 
Two of the furnaces are 14 feet x 22 feet 8 inches hearth dimensions, 
with a capacity of 95 tons ; the third has a 14 feet X 28 feet hearth, 
and a capacity of 110 tons — the fire-boxes being 6 feet 6 inches X 7 
feet. The furnace bottom is constructed of the local silica brick (which 
is claimed to be the finest in the world) laid down in four beds, the three 
lower being each 12 inches thick, whilst the working bed is constructed 
of 20-inch bricks ; brick being found to be better than sand in thift 
class of work. The bottom is curved to a depth of 2 feet. These 
furnaces are, in consequence of their different function, constructed 



— T^ 


on somewhat different principles to the reverberatory smelting 

OA\ing to the high conductivity of copper, and to the fact that the 
functions of the furnace are either largely as a medium for simple 
fusion or as a receptacle for molten metal, and further, that but 
little slag is produced, that no settling and separation of the fluid 
materials are required, and that there is no danger of dusting-losses, 
the furnace may conveniently be built with a deep hearth which need 
not be of very considerable length. The main requirements are 
refractoriness of the building materials, particularly careful construc- 
tion so as to avoid breakouts, and very strong bracing indeed on 
account of the deep and heavy bath of material which is carried on 
the furnace hearth. 

Operations — {a) Oxidation Stage. — The furnace is loosely filled 
with scrap copper which has accumulated round the works (8 to 12^ 
tons), and converter metal (of composition say about 98-3 per cent, 
copper) is then poured in at the side door from ladles bringing it in 
quantities of about 5 tons at a time, as teemed from the converters. 
When the furnace is about half -filled, a blast of air at 90 lbs. pressure 
is injected through the metal by means of iron pipes, which at this^ 
stage just dip below the surface. These pipes are gradually eaten 
away by oxidation and slagging action, but as the end wears down, 
the pipe is pushed further in. The function of this air blast is to- 
supply oxygen for the purpose of acting upon the small quantities 
of oxidisable impurity which remain in the metal after bessemerising, 
and which consist chiefly of iron and sulphur, in addition to the small 
quantities of metalloids. The oxygen partly acts directly on these 
constituents, but as already indicated, the scouring action is to a 
great extent performed by copper oxide which is produced and which 
is itself a powerful oxidising agent. The iron appears to be one of 
the first elements to be removed, and then a little sulphur, but this is 
chiefly eliminated after the iron has been oxidised. The interaction 
between the copper oxide and the sulphides liberates metallic copper 
and yields SOg, which bubbles up through the metal and gives to it 
an appearance of " boiling,'' by which name this stage is known. 
Too rapid an oxidation during the early stages is dangerous if much 
sulphur be present, owing to the evolution of sulphur-dioxide 
isuming a degree of explosive vigour. Up to this point, the oxygen 
las been utihsed in removing iron, sulphur, etc., which are eliminated 
oxides, so that but little of the oxygen is retained in the metal, 
mt after the boiling stage is passed, oxygen is actually absorbed, 
the copper now becoming oxidised, and the oxygen contents of the 


metal rapidly increase. As in the analogous instance of steel bessem- 
-erising, it appears essential to introduce some excess of oxygen into 
the metal in order to ensure the complete removal of the oxidisable 
impurities, so in copper-refining, an excess which amounts to about 
0-7 per cent, of oxygen (equivalent to about 6 per cent, of Cu.,0) must 
be introduced. 

In the refining practice as conducted by the Welsh process, much 
-of this aeration took place during the slow melting down of the crude 
blister copper, and subsequently during the flapping operations with 
the rabble ; but the use of the air-blast hastens this oxidation con- 
siderably, especially as the metal is now often directly poured into 
the furnace in a molten condition, so that oxidation during melting is 
not possible. It is essential to defer this final oxidation and elimina- 
tion until it can be conducted at the refining furnaces rather than to 
attempt it in the converter, since the refining furnace allows of the 
operation being performed much more gradually and under better 
-control, whereas if conducted in the converter, the necessarily vigorous 
action would occasion unduly heavy losses of copper in the slag and 
probably excessive oxidation of the metal. 

During the aeration, the furnace contents are continually added 
to, by additional ladles-full of metal, and usually by the time the 
furnace is filled the air-blast has oxidised most of the impurities from 
the metal. These have entered the slag, and the copper has become 
" dry,'' owing to the necessary super-aeration. If this stage has not 
been reached, the sample often shows " sprouting " (also known as 
" spewing " or '' throwing a worm "), which is caused by the escape 
of SO2, and indicates that all the sulphur has not been eliminated. 
In that case the blowing is continued until small samples ladled out 
from the bath exhibit the characteristics of dry copper, viz.: the 
depression down the middle line of the ingot, brittleness of the metal, 
and a purplish brick-like fracture. 

These preliminary operations may occupy some three or four hours 
or more. 

(h) Poling and Bringing to Pitch. — The oxidation having proceeded 
to the required stage, the highly cupriferous slag is skimmed off, after 
being first thickened with ashes from the fire-grate, and the poling 
of the metal is then commenced. This operation is conducted by 
immersing poles of timber, three to six at a time, in the metal, holding 
them well under the surface and pushing them further in as the ends 
burn away. It is essential that the timber should be green and not 
dry, and preferably it should be hard wood, such as birch, beech, or 
oak. The poles are usually as long as possible, and are from 6 to 8 


inches thick at the butt end. The function of the wood, particularly 
during the early stages is, to a great extent, mechanical, and any 
chemical changes effected are by indirect action. 

The poHng operation really consists of two stages, the first of which 
is the final ehmination of SO.3 retained by the metal, and the last, the 
actual reduction of the excess oxide and the *' bringing of the metal 
to pitch.'' 

The green timber, when inserted into the copper, liberates large 
amounts of moisture and reducing gases which agitate the bath 
considerably and ' shake " the gas out of the metal more or less 
mechanically, replacing, at the same time, some SO., by CO and hydro- 
carbons which copper possesses the power of absorbing. When the 
.S0.> has been satisfactorily eliminated, the reduction stage is arrived 
at. and this is conducted in a manner similar to the familiar poling 
operation of the Welsh process. The surface of the bath is com- 
pletely covered over with a layer of coke, anthracite, or charcoal, 
and more poles are inserted. The exact mechanism of the operation 
has not yet been definitely traced, but the action of the wood at this 
stage is partly of a mechanical and partly of a chemical nature. The 
reducing gases liberated by the charring and destructive distillation 
of the wood have themselves a reducing action on the oxides which 
are dissolved in the dry copper, but an important feature of the 
action of these gases is the agitation and splashing which they occasion, 
thus bringing the molten metal into close and vigorous contact with 
the layer of reducing carbonaceous material maintained upon the 
surface of the bath. 

Poles are inserted usually two or three at a time, and samples are 
constantly taken and examined for surface indications and for fracture. 
This preliminary refining operation usually has for its chief object 
the preparation of a fairly pure metal which will yield a sound, clean, 
and even anode casting, and which is not required at this stage to 
pass the rigid mechanical tests essential for the market product. In 
this case it is therefore usual to carry the poling operation only to 
such a degree that the samples ladled out and cast into small ingots 
soHdify with the even, smooth surface desired and which is charac- 
teristic of "tough-pitch copper" — irrespective of any special mechanical 
properties of the metal. If the test is satisfactory, the metal is ready 
for casting. 

The poling occupies some hours, and usually from 40 to 50 poles 
of wood are used up before the metal is in a suitable condition for 
casting. During these operations the coal fire in the grate is manipu- 
lated in a manner best suited to the various stages of the process ; 

2 24 


there may thus be an oxidising flame during the early part of the 
refining, but the flame must be of a reducing character whilst poling 
is in progress. 

Casting. — Until comparatively recent years, the size of the refining 
furnace has been necessarily limited to small dimensions, owing to 
the difficulty in emptying the furnace of large charges. The practice, 
as conducted hitherto, has been based on the familiar method of the 
old Welsh process, viz., that of ladling out the metal by small hand 
ladles. This involves so much hand labour, and requires such a long 
period of time for its operation as to make practically impossible any 
attempt to deal with large quantities of metal, or to lead to any con- 
siderable increase in furnace capacity. The chief difficulties to be 
overcome when operating on large charges of copper by this hand- 
lading method are those of maintaining the metal at the correct pitch 
during the lengthy period of ladling ; whilst the large amount of time 
during which the finished copper has to remain within prevents the 
furnace being used for its chief purpose, that of refining more metal. 

The method of hand ladling was employed for so many years 
on account of the difficulties of controlling the stream of metal and 
of tapping the furnace in the usual way — i.e.., through a tap-hole at 
the lowest point of the bath. These difficulties were due to the very 
high working temperatures, to the great weight of metal behind the 
stream, which forced it out under great pressure, and to the high 
melting point, conductivity, and tendency to chill of the copper, which 
was apt to cause setting of metal in the tap-hole, and led to the latter 
becoming rapidly closed up and useless. Regulation of the stream 
of metal to a gentle flow was impossible under such conditions. 

With the introduction of casting machines by Walker, and the 
improvements in the methods of tapping by the adoption of the vertical 
tapping-slot, these difficulties have been removed, and the casting 
of 100 tons of metal from one of the modern large casting furnaces 
presents, to experienced workers, little practical difficulty. 

The modern casting machine brings a series of moulds continually 
under the supply of metal which issues from a large ladle fed 
continuously from the tapping-slot of the furnace. 

The method of tapping now used is to allow the copper to gently 
run out of the furnace, by gradually lowering the level of a temporary 
retaining wall which is constructed in a narrow vertical slot in the 
tapping side of the furnace. 

This slot, which is /^-shaped in plan, extends from the lowest 
point of the hearth to well above the highest possible level of the liquid 
metal. It is about 3 feet high and 4 J inches wide, and whilst the 



furnace is working it is kept rammed with a mixture of loam and 
anthracite, this filling being supported by a series of short transverse 
bars, 16 inches long and 1 inch square in section, which are set 3 
inches apart and rest upon lugs fixed to the iron plates which 
strengthen the furnace-wall. During the operation of casting, this 

Fig. 70. — Indicating Tilting and Pouring Mechanism for Ladle of Casting 
and Refining Furnace. 

hard filling can be readily cut away as required and the level of 
the dam thus gradually lowered at will, permitting the gentle and 
ontinuous overflow of the molten metal. The stream is also regulated 
v inserting a pole of wood in the opening, should the flow become 
>o rapid, and by this means it is kept under absolute control. The 
lolten metal flows along a spout which feeds a small suspended ladle 
f about 800 lbs. capacity, the supply being so regulated that this 



ladle is filled sufficiently slowly as not to get ahead of the moulds] 
The ladle is supported hydraulically, and is pivoted so that it cai 
be brought forward and tilted for pouring, and then lowered am 
moved a slight distance backwards, to allow the next mould to come 
into position. 

On tilting the ladle, the metal flows gently and without splashing 
through a three-hole grid in the front — which keeps back slag or cinders-^ 
and runs into the mould, which is rapidly filled. In order to prevent 
the metal overflowing in the mould, and also to rapidly cool that 
portion which forms the lugs of the anode-plate, a hollow water-cooled 
block 2 feet 6 inches long and of 6 inches square section, situated 
opposite the ladle, is brought forward hydraulically into such a 
position that it rests on the mould just against the edge of the lugs. 

At many smelters the circular form of anode casting machine 
introduced by A. L. Walker is employed. This apparatus consists of 
a horizontal wheel which can be rotated slowly, carrying a series of 
arms at the end of which the moulds are supported, so that they form 
a broken ring. By the rotation of the machine, one mould after another 
can be brought under the ladle and filled. The moulds are pivotted, 
so as to allow of tilting, and when the metal has set, the ingot is thus 
dropped into a cold water bosh, whence it is carried to the yards 
by a conveyor. At Anaconda, the casting machine consists of a 
series of moulds carried on a platform conveyor which is operated 
hydraulically — the moulds are attached by bolting them on to the 
belt through lugs fixed underneath. The moulds are constructed of 
1 inch cast iron, and allow of the production of anode ingots 2 feet 
6 inches x 3 feet by 2 inches thick, provided with lugs at the corners 
of one end for the purpose of supporting the plates in the tanks. 

Each mould holds about 560 lbs. of metal, and when the anode 
has been cast, the ladle is dropped back into position and the mould 
is moved forward by means of the conveyor belt. After traversing a 
distance equal to three times its own length, the ingot becomes fairly 
solid, and at a point corresponding to this position the conveyor 
base inclines slightly upwards. The cake is sprayed gently during 
its passage over a distance of about 8 feet, the conveyor belt then 
passes over a pulley-wheel, and when in a vertical position, the anode 
is forced out of the mould by a crowbar and falls into a water bosh 
from which it is carried by another conveyor on to a platform. Hen 
it is wheeled to stacks, examined for flaws, and weighed. Sample anodes 
are placed on one side, and the others are packed for shipment to th« 
Eastern refineries. 

The furnace deals with one charge (usually of 100 tons, bu 

[1 fare pajfc "220. 

Fig. 71. — Walker's Anode Casting Machine. 

Fig. 72.— General View of Tank-room of Electrolytic Refinery, Perth Aniboy, N.J. 


occasionally much more) per eight-hour shift, and the casting machine 
yields 25 tons of anodes per hour. 

Samples weighing from 4 to 6 ozs. are taken three times per shift 
from the stream of copper running into the moulds, by batting the 
metal into water with a wooden paddle. This method checks very well 
with drillings taken from the anode plates, the chief discrepancy feared 
having been with respect to silver contents, owing to the tendency 
of this metal to segregate. The assay of the anode metal at Anaconda 
averages copper 99-3 per cent., silver 80 ozs. per ton, and gold 0-5 oz. 
per ton. 

Electro- Refining. — Electrolytic refining was introduced on a 
•commercial scale by Elkington at Pembray in 1865, and with the 
general adoption of the dynamo for the production of power, dating 
from about 1870, the process was greatly developed. Most of the 
copper now placed on the market has passed through the electro- 
lytic refinery. 

System of Working. — The method of arranging the electrodes in the 
depositing tanks which is usually adopted at the great refineries at the 
present day, is that known as the p arallel or mul tiple system. 

In this method of working, tHe'anodes are aH connected to one 

I pole of the circuit, and the cathodes, situated between them, are all 

•connected to the other. In this way, each tank comprises in reality 

one large anode and one large cathode, and the voltage as measured 

' letween any two neighbouring electrodes will be the same. The system 

hus allows of currents at low voltage being employed, since the voltage 

s a factor of the number of electrodes in series, and in consequence 

1 anger of short circuiting is lessened. This allows of plates being 

ilaced closer together in the tank, with less danger from this source of 


A large number of tanks are employed at the refineries, and 
they are usually arranged in series, the anode plates of one vat being 
onnected to the cathode plates of the neighbouring one, the current 
'hus passing from one vat to the other through the entire system. 

Various other methods of arranging the electrodes have been 

favoured from time to time, and of these the aeries-ay stem is the 

most important, this being still in use at several large refineries, 

^^though it has been generally superseded by the multiple method. 

I^B The plan underlying the series method was that of avoiding the 

l^fcouble and expense of preparing and working with the special cathode 

■Hoeets of pure copper as are necessitated by the multiple system. 

In the series-method, each anode was made to serve as a depositing 

surface for the pure cathode copper produced by the operation, so 


that as impure anode copper was dissolved away on one side of 1 
" anode ''-plate, pure copper was gradually deposited upon the othei 
side. This system appeared, therefore, to have several marked advan-^ 
tages to recommend it, but in practical operation many difficulties in 
working and several serious disadvantages were encountered. The 
chief points in favour of the series-system are — 

(a) Smaller first cost of the installation, particularly in the matter 
of electrical connections, since the multiple system requires heavy 
leads running along each side of the tank, as well as close attention 
to the providing of good contacts, in order to connect all the anodes 
and all the cathodes together with a minimum of current leakage. 
In the series-method, the plates are readily connected one to the other. 

(h) The great saving of the cost of preparation and arranging for 
specially pure cathode plates, this constituting a very important factor 
in the costs of the multiple process. 

(c) The output of metal per vat is greater. 

On the other hand, the disadvantages of the system, except under 
special conditions, are very serious. 

(a) More scrap is produced and requires re-treating, owing to the 
difficulty of separating the new deposit from the remaining portions of 
old anode, which often adhere very firmly. 

(6) Higher voltage through the tanks is required, owing to the 
large number of electrodes in series in the bath. Hence the danger 
of current-leakage and short circuiting is greatly increased, especially 
when impure anodes are used, since they tend to produce conducting 
layers of mud on the bottom of the vat. 

(c) The anode plates have to be made particularly smooth and 
even on the surface, since in order to lessen the voltage required, 
the plates are brought as close together as possible, in consequence of 
which, any excrescences upon the surface greatly increase the danger 
of short circuiting. Prehminary furnace- refining and special straight- 
ening of the anodes are therefore essential in connection with the series 

(d) Special tanks are required, as the protecting lead liner cannot 
be employed, since the danger of current-leakage through it is 
increased, owing to the higher voltages required. Hence special' 
acid-resisting material, such as slate, is necessary, the expense of which 
is considerable. 

(e) The cost of stripping the cathodes is high, and the operatior 
is often difficult. 

(/) The cost of maintaining the plant is greater. 



The special advantages of the multiple system are that — 

(1) It is apphcable to all grades of metal ; 

(2) It permits of either high or low current-density being 
employed ; 

(3) It permits the use of cheaper lead-lined wooden tanks for 
working ; 

(4) Anode plates may be used without previous refining, if so 

It is, however, general to carry out this preliminary refining, which 
yields sounder anodes, keeps the electrolyte purer, and promotes the 
more regular working of the electrodes and electrolyte — although 
-ome smelters still cast the anodes direct from converter metal. 

Summarising, it is found generally that — 

(1) A marked saving is effected in operating the multiple system 
(this has been estimated by Barnett, at as much as $2 per ton of 
refined metal — 8s. 4d.) ; 

(2) A greater efficiency is obtained, the tank efficiency of the 
multiple system being 95 per cent, compared with 90 per cent, for 
the series-system ; 

(3) Less copper is held up in the multiple system, since less anode 
copper is required, under Hke conditions as regards cathode surface y^^ 
and current density. ^ 

Outline of the Process. — The electro-refining industry is a 

highly speciahsed one, and the methods of putting the comparatively 

simple underlying principles into practical operation have assumed 

ireat complexity and diversity in detail, concerning which Ulke has 

ollected and published much valuable information. 

The following details have more particular reference to the multiple 
system of working, as being the most representative of the electro- 
refining methods in general use. 

The outlines of the process constitute the passing of direct electric 
current through tanks containing acidified solutions of copper sulphate, 
' mploying plates of crude copper as anodes, and depositing pure metal 
ipon cathode-plates of specially-refined thin sheets of pure copper. 
The precious metals and most of the impurities of the anode metal are 
Hberated as small insoluble particles which gradually settle to the 
bottom of the tanks in the form of mud, soluble constituents, such 
as iron and zinc, first passing into solution. 

General Conditions — Anodes. — The usual dimensions of the anode- 
plates are 3 feet high by 2 feet 6 inches wide and about 2 inches thick ; 
they are generally cast with lugs, so as to allow of suspension in the 


tanks. The anode metal is usually brought by a preparatory opera- 
tion, to as high a state of purity as is economically practicable — 

(a) In order to obtain smooth and sound electrodes. 

(b) To ensure better working in the tanks. 

(a) The necessity for the employment of solid and even anodes has- 
already been indicated ; it allows of closer suspension of the elec- 
trodes, lessens the liability of sprouting and unevenness on the deposits- 
and the irregular wear and breaking up of the anode-plates before they 
are sufficiently worn away. 

(6) The more free the metal is from iron, sulphur, zinc, nickel, 
etc., the purer remains the electrolyte, since these elements pass into- 
solution at a greater speed than does the copper itself, and, gradually 
concentrating in the tank liquors, render them more and more impure — 
the purity of the metal deposited at the cathode being in consequence 
decreased. The preliminary refining and bringing up to pitch of the 
metal before casting into anodes, as already described, thus has for 
its object the preparation of electrodes in a suitable mechanical as 
well as chemical condition. The copper content is rarely less than 
98 per cent, and is often more than 99 per cent. The gold and silver 
contents are not affected by this preliminary treatment, nor are, 
to any great extent, the proportions of arsenic, antimony, bismuth, 

The size of the anode-plate varies somewhat at different refineries, 
the usual standard dimensions being indicated above ; the size 
depends to a large extent upon the facilities for handling the electrodes 
and on the circuit system operated. There is a tendency at several 
works possessing suitable facilities, to increase the size of the electrodes. 

The Cathodes. — The copper is deposited from the electrolyte upon 
cathode sheets, which are usually thin plates of pure copper corre- 
sponding in size to the anodes. As these sheets cannot be conveniently 
provided with suitable lugs for suspension, they are usually made of 
somewhat greater length than the anodes, so as to allow of bending 
over the cross-conductors ; otherwise they are furnished with metallic 
clips for attachment to these bars. 

These cathode sheets are prepared by depositing layers of pure 
metal upon plates of refined copper of suitable surface dimensions 
and of about J inch thickness. Each side of these plates, which are 
specially smoothed, is first slightly oiled so as to allow of the 
subsequent convenient stripping of the sheet when made, and it is 
then well coated with graphite in order to present a conducting 
surface on which deposition can proceed. The cathode-sheets are 


deposited either in the regular tanks of the refinery or in vats 
specially devoted to the purpose, using in that case, pure electrolyte, 
and working in the usual manner. On attaining a thickness of about 
oV inch, the sheet is stripped off, cleaned and clipped ready for use 
as a cathode. 

The Electrolyte. — The electrolyte is essentially an acid solution of 
copper sulphate. The average proportions are from 15 to 20 per cent, 
of copper sulphate crystals, and from 5 to 10 per cent, of sulphuric 
acid — the usual density of the solution ranging from 1-12 to 1-25. The 
liquid under ordinary conditions of working, remains reasonably pure 
for a considerable time. It tends, however, to decrease in acidity and 
to increase in copper contents, partly owing to the presence of cuprous 
oxide in the metal, which passes into solution independently of the 
indirect transference of metallic copper from anode to cathode. The 
composition of the tank liquors must, therefore, be frequently checked. 
The gold and silver values do not pass into solution under ordinary 
working conditions, and the addition of a small quantity of common- 
salt or of hydrochloric acid to the vat effectually prevents any silver 
from remaining in solution in the liquors. 

A considerable proportion of the arsenic in the tough-pitch anode- 
copper, existing as arsenate, is deposited with the mud residues, it 
being insoluble ajad non-conducting. Arsenic in a reduced condition is, 
howeverTsoIuble, and may gradually concentrate in the liquors and 
contaminate the cathode copper, unless suitable precautions are taken. 
Some of the reduced arsenic, moreover, tends to produce a slimy 
arsenite of copper, which, though insoluble, exists in a colloidal non- 
settling form. Addition of ammonium sulphate to the electrolyte 
prevents this formation, whilst combined aeration and heating pro- 
mote the precipitation of arsenic as insoluble arsenates which settle 
with the tank slimes. 

Some of the antimony and bismuth tend to first pass into 
solution, but for the most part they are precipitated as insoluble basic 
salts. Under suitable conditions with respect to the acidity and 
copper contents of the electrolyte, there is little tendency for 
deposition of these impurities with the copper, but deviation from the 
correct composition is liable to cause contamination of the deposited 
metal. These impurities, when in solution, tend to be oxidised by 
aeration, and this operation greatly encourages their precipitation 
with the mud. 

Iron readily dissolves in the electrolyte, forming soluble ferrous 
sulphate which tends to gradually accumulate in the solution. This 
contamination spoils the quality of the deposited metal, and inter- 


feres with the process of deposition, decreasing the conductivity of the 
bath and thus necessitating higher voltage. Aeration of the solution, 
especially when warmed, leads to the formation of basic ferric 
sulphates which are insoluble, and which therefore accumulate at the 
bottom of the tank. 

Selenium and tellurium, which when present most probably exist 
as insoluble selenides and tellurides of copper or silver, are also 
precipitated, and thus do not find their way into the deposited metal. 

The copper itself is deposited from the electrolyte on to the cathode- 
sheets by the action of the current, whilst at the anodes, the metal 
passes into solution, and the other constituents are either dissolved 
or precipitated. It follows that in an undisturbed solution, the 
liquid near to the cathode becomes gradually impoverished in copper, 
resulting in a decrease in the rate of deposition and necessitating 
greater electrical pressure, whilst in the neighbourhood of the anode, 
the liquor is proportionately stronger in copper and less acid in 
character. Should these conditions continue to any great extent, the 
working of the bath is seriously interfered with, since diffusion 
proceeds too slowly for uniformity to be restored, and in order to 
secure uniform composition of the electrolyte, it must be maintained 
in gentle motion by some system of circulation. This agitation and 
mixing is assisted by the aeration of the bath for the purpose of 
hastening the oxidation and precipitation of several of the impurities — 
this being effected by blowing through the liquid a gentle supply of air. 

Temperature. — The electrical resistance of the solution decreases 
as the temperature rises, and in practice the bath is maintained at 
a uniform temperature of 45° to 50° C. By this means a useful 
increase in conductivity is obtained, the strength of the deposited 
copper being at the same time greatly augmented. 

Electrical Conditions. — The electrical factors which mainly control 
the working of the electrolytic process are those of — 

(a) Current density. 
(6) Voltage. 

(a) Current Density. — The quantity of metal deposited from an 
electrolyte is proportional to the current which passes, and to the 
electro-chemical equivalent of the particular metal. Thus a current of 
one ampere will deposit from copper sulphate solution, M832 
grammes of copper per hour, and the total quantity deposited in any 
given time is determined by the product of the current, the time, 
and this electro-chemical equivalent (which is determined experi- 


In practical operation, a factor which is amongst the most important 
of those governing the working of a plant, is the current density, or 
current per unit of area of depositing surface, since from this factor 
the rate of deposition upon the cathode plates is determined, and from 
it the power requirements, accommodation, etc., for the plant are fixed. 
The current density is subject to wide variation, but, as a general rule, 
it ranges from 8 to 18 amperes per square foot of plate-area. Its value 
is largely dependent upon the speed of working, the cost of power, 
and the composition of the anode metal, the electrolyte and the 
desired product, etc. 

In general, high current- density possesses many advantages, 
resulting from the fact that it occasions a more rapid deposition. 
It causes a proportionately greater output, consequently the stock 
of metal held back in the tanks is reduced, and hence there is less 
capital locked up in the form of metal undergoing treatment, and less 
plant and accommodation are required for the same output. 

The current density permissible is, however, limited by the com- 
position of the electrodes and the solution. High current-density 
causes rapid dissolution of the anodes, and if the plates are not 
particularly free from impurity, the electrolyte rapidly becomes 
contaminated, since its dissolving power on the impurities becomes 
greater with increased electrolytic action, and this affords less 
opportunity for the precipitation and settling of the injurious con- 
stituents. In consequence, the cathode copper is contaminated 
through this mechanical inclusion of impurities, whilst electro- 
deposition of some of these materials may also be encouraged. The 
presence of much silver in the anodes causes the rapid breaking-up of 
the plates, especially if the current density be high, and thus the 
separation of the values in the slimes is not so efficiently managed. 
With high values in the anode copper, it is necessary to reduce the 
current density to 8 or 10 amperes per square foot, whereas with purer 
metal a density of as much as 16 to 20 may be conveniently employed. 

(b) Voltage. — Electrical pressure is required in order to force the 
depositing current through the electrolyte against the resistances in 
the circuit. The voltage required depends upon the current density, 
the composition and temperature of the electrolyte, the composition 
•of the anodes, and also upon the general conditions of working. These 
being constant, the voltage necessary is largely a factor of the number 
of electrodes in series in the tank and of the distance apart of the 
plates. Under ordinary circumstances this voltage varies from 0- 1 to 
About 0-3 volt. High voltage is to be avoided, owing to the danger 
of short-circuiting, especially in cases where the accumulation of mud 



in the tanks, or the impregnation of metallic salts in the tank walls, 
or the growth of excrescences upon the plates, lead to the passage of 
the current through these conductors rather than through the electro- 
lyte solution itself. Short circuiting naturally diminishes the output 
of the plant. 

These electrical factors which form the basis of the power require- 
ments of the refinery, call for careful observation during the progress^ 
of the operations in order to ensure successful working and a high 
efficiency of the plant. 

The current from the dynamos is brought by heavy leads and is 
distributed through the sets of tanks in the manner best suited to 
the installation. At one of the newer works, dynamos producing 
about 6,000 amperes at 120 to 150 volts, supply sufficient current tO" 
operate a set of 400 vats working on the lines just indicated. 









Glass „ 



Fig. 73. — Indicating Methods of Suspending and Connecting Electrodes 
(Perth Araboy, N.J.). 

The Depositing Tanks. — The tanks are usually constructed of 
wood, such as strong pitch pine, and they are lead-lined. The cross- 
section is usually such as will allow a space of about 3 inches between 
the edges of the electrode and the wall, and a 6-inch space from the 
tank-bottom to the lower edge of the plate. The length of the vats 
varies considerably, according to the desired output and to con- 
venience of working — 10 to 15 feet being average dimensions. This 
size of vat will hold 15 to 25 anodes together with a corresponding 
number of cathodes (16 to 26). The tanks are arranged across the 
building in a number of rows which are usually stepped down in- 
stages of about 2 to 3 inches each, so as to assist as much as possible 
the circulation of the electrolyte through the system by gravity, and. 
the vats are set in pairs with aisle-Avays between (see Fig. 72). 

Leads run along each side of the tank, the current being conveyed- 




to all the anodes at once by resting one lug of each plate upon the 
lead which runs along one side of the tank, carefully insulating the 
other lug from the conductor situated at the opposite side, this being 
used for connecting up the cathodes. The cathode-sheets are sus- 
pended from metallic cross-bars, which rest upon their own conducting 
lead, and are carefully insulated from the anode lead at the other 
side. The solutions are heated by means of steam coils. 

Distribution of the Electrolyte. — The necessary circulation of the 
electrolyte is eflFected as much as possible by the natural action of 
gravity. The tanks of the top row in the deposi ting-house receive a 
constant supply of fresh solution from upper distributing vats, 



Cross Section qT Tank. Showing Tube and Base 


'ubber tfushin 

Tubing [ 


4-h— 6"'-i- 


Fig. 74.— Indicating Connections for Circulation of Electrolyte (Barnett). 

I whilst old electrolyte is drawn off from near the bottom of the 
tanks, and flows over to those on the next and lower level. Fresh 
solution thus enters at the top of the tank, old solution is drawn off 
from below, and thus a uniform density and composition is maintained. 
From the tanks situated at the lowest level, the solution passes to a 
well, and from there is pumped up to the store-tanks or, when 
necessary, to the purif3dng tanks; air-pressure pumps being often 
employed for this work. 

In course of time— and under the modem system of working with 
moderately pure anodes, this period is of considerable duration — the 
gradual accumulation in the electrolyte, of the small quantities of 


impurity which are dissolved from the anode, may render the Hquid 
so impure, that a danger arises of contamination of the cathode 
copper to such a degree that it becomes unfit for conductivity work. 
It then becomes necessary to purify the solution. In present-day 
practice, this continued accumulation of impurity in the electrolyte is 
prevented by continuously withdrawing, for separate purification, a 
certain proportion of the electrolyte from the circuit — replacing it by a 
fresh supply of pure solution from the store-tanks. Constant regenera- 
tion, purification and circulation are thus effected, whilst uniform 
composition is maintained. 

After considerable use, the electrolyte solution gradually tends to 
increase in copper contents, and the first stages in the scheme of 
treatment for the old solution is to recover this excess of copper, 
which is effected in tanks known as "liberating tanks.'* These are 
similar in general features to the refining vats, except that lead plates 
are employed instead of the copper anodes, so that the excess metal is 
deposited without any addition of copper being made to the solution, 
from the anodes. In due course, the desired composition in the 
electrolyte is once more attained. 

When the solutions have become too impure for further use in the 
tanks, the bulk of the copper sulphate is recovered by evaporation 
in large pans, followed by crystallisation in somewhat shallow vats of 
large dimensions. The crude blue vitriol is further purified by re- 
peated crystallisation, and any copper which still remains in the 
solution is then precipitated on scrap iron, the cement copper being 
worked through the furnaces again. Excess acid is also often 
recovered on further evaporation of the Hquors, and is employed in 
the subsequent treatment of the sHmes. 

Working. — In the large modern refineries, the anodes are carried to 
and from the tank-house by cars, and at the tank- room are suspended 
from frames which are conveyed over the baths by means of overhead 
•electrical cranes of about 10 tons lifting capacity. These rectangular 
frames correspond in size to the dimensions of the tanks, and are 
constructed of steel girders. Under the longer sides of this frame a 
series of hooks project, upon which the lugs of the anodes rest, and 
the hooks are placed at distances corresponding to the eventual 
position of the plates in the tank, so that the whole series of anodes 
can be dumped into position at one operation. 

The cathodes are placed in a second rack, and likewise brought into 
position, between the anodes. The solution is then turned into the 
tank, the current started, and the refining proceeds, with a steady 
ilow of liquid circulating through the system. The operations of 



changing electrodes, cleaning and reloading occupy about one houiv 
and, but for this manipulation, the process under normal working is 
continuous. In ordinary practice, about 20 to 25 lbs. of copper are 
deposited daily on each cathode. Constant examination is made as 
to the electrical conditions, and the composition, temperature, and 
density of the solutions. 

The anodes usually remain in the bath for a period of about 
six weeks, and they are then removed from the tank, scrubbed, and 
sent back to the furnaces to be re-melted and re-cast into fresh 
anodes, the quantity of such anode scrap under good working 
conditions amounting to about 9 or 10 per cent, of the original metal. 

Fig. 75. — Tank-house, showing Anode Crane (Ulke). 

The cathodes remain in the tanks for about one week, by which 
time a deposit of from 150 to 170 lbs. of pure metal has been obtained 
upon each. The practice of frequently replacing the cathodes possesses, 
among other advantages, those of maintaining a more even current 
density over the plates, of preventing the growth of excrescences and 
the irregular dissolution of the anodes, and of lessening the danger 
of breakdown of the somewhat slenderly suspended cathodes, by 
putting less weight on the supports. The removal in one operation of the 
entire batch of cathodes from the bath is effected by means of the 
suspended hook-frame, as employed in charging. The plates are 
rinsed, the top edges are cut off and returned with the anode scrap, 
whilst the pure electrolytic copper passes to the refining and casting 
furnaces, where it is prepared for the market. 


Collection of the Slimes. — Depending upon the working conditions 
of the refinery, but usually at intervals of three months, the precipi- 
tated slimes are collected and the tanks are cleaned out. The quantity 
of shme deposited is generally not very large, from 15 to 25 lbs. per 
tank being a not unusual yield. The current and the supply of 
solution are cut off, the plates removed, the contents of the tank 
Billowed to settle, the liquid siphoned off to within about 6 inches 
of the bottom, and the residues are swilled out through a trap at the 
bottom of the tank. The sludge passes through a sieve that separates 
the lumps of anode copper which have broken off and fallen to the 
bottom of the tank, the slime then passes to the special refinery for 
treatment. The processes adopted for recovering the gold and silver 
from this residue are highly specialised, and belong properly to the 
technology of refining of the precious metals. 

Modifications of Electrolytic Refining. — Great success has not yet 
attended the attempts which have been made to employ copper matte 
in the form of anodes in electrolytic refining processes, and the method 
is not in operation at any of the great modern works. Marchese, 
Hoepfner, Siemens-Halske, Keith, and others have introduced pro- 
cesses, but their practical operation is attended with very great 
difficulty and but little commercial success. Matte is exceedingly 
brittle and it readily breaks up, it is a bad conductor and 
necessitates the use of high voltage, the solutions become very foul, 
^nd the processes require very special apparatus and equipment. 

Methods for the production, by electrolytic processes, of pure copper 
in forms ready for service, such as wires or tubes, have been introduced 
successfully by a number of workers, including Elmore, Thomerson, 
-and Cowper-Coles. Several of these methods are now in apparently 
successful commercial operation, and the published results of the 
working of the processes and of tests on the deposited materials offer 
considerable promise for their future industrial application for special 
purposes, if not for general use. The attaining of the necessary 
•compactness, toughness, and strength of the metallic product is aided 
by the employment of pressure during deposition, as by burnishers, 
or by very rapid rotation of the depositing surfaces in the solutions. 
Details of these processes and products may be found from the 
references subsequently given. 

Bringing up to Pitch and Casting the Merchant Copper, — 
The final stages in the smelting process from ore to market-metal are 
those of ''fining,'' toughening, and casting the cathode copper, the 
object of these operations being to impart to the metal the chemical 
composition and mechanical and physical properties which are 


required in order to fit it for the market, and also to prepare it into a 
suitable form for service. In addition to cathode copper, other forms 
of the metal, if of suitable composition, are also treated with this 

For conductivity copper, however, these final operations are con- 
ducted on metal from which practically all the impurities have been 
removed, but which is not sufficiently tough and homogeneous or which 
is not in a suitable shape for immediate industrial use. The toughening 
operation consists almost entirely of adjusting the percentage of 
oxides in the metal, partly in order to overcome the influence of any 
traces of injurious impurity that might remain, but mainly to 
exercise the functions previously indicated, of imparting by its more 
or less direct action upon the metal, a definite toughening and 
strengthening effect. The mechanism of the action is not perfectly 
understood, but the recent work referred to in Lecture II., p. 28, 
affords useful evidence as to its possible mode of action. 

The actual refining operation and the furnace employed for the 
process are exactly similar to those used in preparing the metal to 
ensure the casting of sound ingots, as already described. The opera- 
tions consist of a preliminary aeration, by means of which any 
oxidisable impurity still remaining in the metal is oxidised out, 
mainly through the action of copper oxide which is formed during the 
process in some considerable excess. 

After the copper has become " dry " or over-oxidised, which condi- 
tion is characterised by brittleness, depressed surface, and brick-like 
purple-red fracture of the metal, it is reduced by poling and timbering 
operations to a definite point, viz. : until a sample ingot of the metal 
indicates a maximum of toughness, accompanied by level surface and 
bright salmon-coloured silky fracture — it is then of " tough-pitch " 

The furnace employed for the refining has already been described. 
One of the main features in which it differs from the ordinary modern 
reverberatory smelting-furnace is that owing to the exceedingly high 
heat-conducting power of metallic copper, and to the absence of an 
insulating layer of non-conducting slag, there is little danger of much 
chilling action occurring on the hearth of the furnace ; the temperature 
may, indeed, often become too high rather than too low. In conse- 
quence, it is not so usual to construct the furnace with a very massive 
hearth foundation as for smelting, but to build it upon a vault or 
upon a series of piers. With this type of foundation, the very con- 
siderable, but practically unavoidable, absorption of metal in the 
hearth-material is reduced to a minimum. It is usual to work a 



charge consisting of scrap and oxide in the furnace before the regular 
smelting campaign begins in order to " season '' the hearth. This 
procedure allows the primary absorption of copper by the hearth- 
material, and assists its consolidation, whilst the action of the oxide 
promotes a surface glazing which lessens the tendency for further 
absorption of copper, and gives a good surface to the working bed. 
As has been already stated, the hearth is generally built of brickwork 
rather than of sand. The furnace is constructed to hold from 80 up 
to 200 tons of metal. The method of working differs mainly from that 
previously described, in that instead of pouring molten copper into 
the furnace, as is usual with converter-metal, the cathode plates 
must be charged in a different manner. 

In order to deal with such a large quantity of charge in this bulky 
form, without occupying so much time as to make the whole operation 
too protracted, it is usual to employ some form of charging machine 
rather than to use hand labour for the operation. In some cases a 
small melting furnace is employed solely for the purpose of preparing 
the metal in a molten form for feeding into the refining furnace. The 
type of cathode-charger most used is very similar in operation to the 
Welman charger for steel furnaces, and by its means, 100 tons of 
material can be charged per hour. 

Operation. — The refining and toughening process is conducted in 
the six stages of : — 

(a) Charging. 
(6) Melting. 

(c) Skimming. 

(d) Oxidation, by aeration, 
(c) Reduction, by poling. 
(/) Casting. 

(a) The charging is sometimes conducted in stages, this being 
indeed unavoidable when very large quantities of material are worked 
with, the bulk of which, when solid, would more than fill the furnace. 
Two-thirds or three-quarters of the material may be put in at first and 
just melted down slowly, after which the remainder is added. 

Owing to the not infrequent presence of sulphur in the furnace 
coals, and to its ready affinity for copper, resulting in undesirable 
consequences for the commercial metal, contamination by this element 
is usually prevented, as much as possible, by giving to the cathodes a 
wash of lime previous to charging. 

ip) The melting is generally conducted somewhat slowly, so as 
to allow some oxidation of the metal during this stage, which may 


occupy some twelve hours. Skimming of slag as it forms, and 
subsequent blowing of the copper towards the end of the melting 
stage are frequently resorted to. 

{c) The slag which accumulates, sometimes in considerable quantity, 
is skimmed off as occasion requires. When converter metal of such 
purity as not to need electrolytic refining is treated directly in the 
furnace, much of this slag is converter-slag introduced from the ladle, 
and requires to be skimmed off at an early stage. In the usual pro- 
cess of melting cathode-copper, slag is produced from the last 
traces of iron which may have remained in the metal. In order to 
render it sufficiently viscid to be pulled out by the skimmer, ashes 
from the fire-grate are thrown upon and rabbled into the slag. This 
skimming may continue for some time, and a very rich coppery slag 
is pulled off, from which the metal values are subsequently recovered. 

(d) The oxidation of the small quantities of impurity still remaining 
in the metal is completed by the operation of airing, as already de- 
scribed, and the action is continued in order to produce a small excess 
of oxide until the copper is " dry.'" The time occupied for this airing 
is now not very protracted, since most of the impurities have been 
previously removed from the metal. 

(e) The copper is then brought up to pitch by " poling " in the 
manner previously indicated, except that* at this final stage, the testing 
of the metal and the adjusting of the oxygen proportion are conducted 
with much greater precision than was necessary for the simple produc- 
tion of the sound anode plates. In the present instance, the character 
of the metal and its value as a commercial article largely depend 
upon the care and accuracy with which the correct " pitch " is reached 
and is maintained in the bath during the entire period of casting of 
the metal. The poling for the " shaking out '* of the gases is rarely 
necessary with cathode metal, and the addition of the cover of 
carbonaceous material for the purpose of effecting the reduction of the 
oxides to the desired extent, is made either at the commencement of 
poUng or else shortly afterwards. After some time, a series of small 
samples is taken at intervals, by means of ladles, and the surface of 
the ingot is examined. The depression characteristic of dry copper 
gradually becomes less marked, the brick-like fracture appears finer and 
finer until it becomes silky, whilst the colour eventually turns to a 
very dehcate salmon-pink. Meanwhile the mechanical properties have 
gradually improved, signs of brittleness disappear, and somewhat 
larger samples of the metal, which are now taken and tested, are 
characterised by a very marked toughness and strength. This is 
the moment at which the poling must cease. The residual copper- 



oxide has now reached the proportion which was necessary for the 
imparting of the best mechanical properties, and the metal is tough- 
pitch. The skill of the workman is now exercised to the highest 
degree, in maintaining the metal in this condition during the whole 
of the subsequent casting period. Oxidation must be avoided in 
order to prevent a reversion to dry copper, whilst any further reducing 
action removes some of the necessary oxide, and results in " over- 
poling." The metal would then become brittle again, coarsely fibrous 
and possibly somewhat spongy in fracture and very pale in colour, 
whilst in setting it would show a ridge upon the surface. In that 
case it would be necessary to " air " the metal again until it became 
dry, and then to pole it back to the " tough-pitch '' stage. 

The copper, when of correct pitch, is therefore removed from the 
furnace and cast at once; this being readily conducted through the 
tapping slot, the level of which is gradually lowered. The metal then 
flows down the spout to the ladle, and is poured into the moulds 
attached to some form of mechanical casting machine ; the ingots 
being finally dropped into a water-bosh, weighed, sampled, and 
stacked, and are then in a condition ready for the market. 

Phosphorus is sometimes employed for giving soundness to the 
castings, being added to the bath in small quantities in the form of 
phosphor-copper containing about 10 per cent, of the non-metal. 
Although very little of this phosphorus is retained by the metal, being 
mostly eliminated as oxide, special caution is required in emplopng 
it for high-grade conductivity copper, since the effect of very small 
quantities has a deleterious influence upon the conducting properties. 

Silicon also is used for a similar purpose, and causes a considerable 
increase in toughness. 

(/) When intended for conductivity work, the metal is cast into 
the form of '' wire-bars "' of very varied shape and size, according to 
requirements ; thus the 100-lb. bars are about 3 feet long by about 
3 inches square section, the 500-lb. bars 7 feet long by about 4| inches 
square. Furnace samples weighing about 1 lb. are drawn down 
gradually to about |-inch wire, and are tested for conductivity, as 
well as for strength and toughness, occasional analysis being also 
undertaken, whilst samples of the wire-bars in market form are 
similarly examined. 

[To Jacr t>»y<' U'i- 

^: ' 

Fig. 76. — Microstructure of Commercial Copper containing O-xygen (Hofnian). 

a. Calumet and Hecla copper after b. Calumet and Hecla "dry " copper 

60 minutes' poling. before poling. 

22 per cent, oxygen = I 98 per cent. CU2O. 64 per cent, oxygen —5 •76 per cent, Cu^O. 

Compare with Fig. (5, p. 28. 

(lii/ penninHioii oft.ha American J nxtitution of Minittg Kngineert.) 



Ulke, T., " Modern Electrolytic Copper Refining." (With complete Bibliography). 
Peters, E. D., "Modern Copper Smelting" (1905). (Chapter xviii. by M. Barnett.). 
Mineral Industry. Annual Review. 
Cowper Coles, S., "An Electrolytic Process for the Production of Copper Wire." Proc. 

Rham. Met. Soc, 1908-9, p. 5. 
Schnabel and Louis, "Handbook of Metallurgy," pp. 327-369. 
Wraith, W., "Sampling Copper Anodes at Anaconda." Tram, Amer. List. Min. Eng., 

March, 1910. 
■"The De Lamar Electro-Refinery at Chrome, New Jersey." E7ig. and Min. Jotirn., Jan. 

13, 1906, p. 73. 
Flinn, F. B., "Electrolytic Copper." The' Metal Indnstri/, April, 1910, p. 112, vol. ii.. 

No. 3. 
<Treenawalt, W. E., " The Greenawalt Electrolytic Process." Eng. and Min. Joum.y Nov. 

26, 1910, p. 1062. 

See also References to Johnson (p. 34), Hofman and others (p. 50), 
Keller (pp. 50, 215), and Peters (p. 80). 



Accretions in the blast furnace, 55, 124, 

130, 158. 
Acidity of electrolyte, 231, 236. 
Acid-lined converters, 193-213. 
Acid-making from blast-fiirnace gas, 140, 146, 

167, 181, 186, 188-191. 
Acid-making from roaster gas, 63, 68. 
Acid-plant, 190, 191. 
Acid- silicates, 148. 
Acids and copper, 33. 
Addicks, L., 23, 34. 

Advantages of basic linings, 202, 204, 213, 
„ blast furnaces, 114. 

„ Briickner roasters, 72. 

„ forced draft, 85. 

„ high current density, 233. 

„ large grates, 84. 

„ ,, reverberatories, 91. 

„ long blast furnaces, 129. 

„ matte-pool, 91. 

„ MacDougal roasters, 75, 76. 

„ multiple system, 229. 

„ pyritic smelting, 177. 

„ roasting, 62, 67. 

„ sectioning, 124. 

„ series system, 228. 

„ upright converter, 195. 

„ water- jacketing, 123, 124. 

Aiiration in electro-refining, 231, 232, 

furnace refining, 222, 239-242. 
Africa, copper output, 15. 
Agglomeration of fines, 47, 51, 55-57, 110-112. 
Air and copper, 33. 

for furnace refining, 221, 222, 241. 
in roasting, 66. 

-admission in reverberatory work, 86, 
-holes in reverberatory furnace, 85, 86, 101 . 
-space under blast furnaces, 136. 
-supply for blast furnaces, 141-145, 188. 
converters, 195-198, 213. 
„ pyritic smelting, 173-188. 
„ reverberatory work, 85, 92, 
.\iring, 41. 

Alabaster, R. C, 191. 
.\laska copper output, 15, 
Alchemists, 2. 
Allen roaster, 70, 
Alloys, 18, 21, 34, 40, 42. 

Alumina in slags, 147, 150, 191. 
Aluminium and copper, 21, 23. 
Amalgamated Copper Company, 13. 
American copper mining, 7. 

„ production, 15. 

Ammonium sulphate in electro-refining, 231. 
Anaconda, 9, 52, 80, 191, 220, 243. 

„ blast-furnace practice, 120, 190. 

blast furnaces, 125-128, 166, 191. 
„ briquetting plant, 57. 

„ Briickners at, 72. 

„ Casting at, 226, 

„ charge calculations, 150. 

„ „ -cars, 154, 155. 

„ converter practice, 194, 201, 207, 

converters, 194, 197, 207, 216. 
„ copper, 227. 

„ costs of production, 14. 

„ refining, 219. 

„ reverberatory practice, 82, 85, 89, 

91, 96-104. 
„ roasting, 74-79, 

„ sampling, 47, 48. 

„ smelting scheme, 54, 

„ wet concentration, 53. 

Analysis of copper, 20, 44, 50, 227, 

,, costs of oil-fired reverberatories, 

81, 107, 108. 
„ „ roasting, 79. 

Annealing of copper, 28, 31, 32. 
Anodes, 217, 219, 226, 229-238. 
Anode-casting machines, 226, 

„ -copper, 54, 219, 230, 233, 243. 
„ -scrap, 237, 
Antimony and copper, 20, 23-25, 30-32, 44, 
207, 217, 230, 
„ in copper matte, 37, 207, 217. 

in electro-refining, 217, 230, 231. 
Apparatus for roasting, <56, ei /tet/. (see 

RoaMing fiirruwes). 
Appearance of copper- mattes, 38, 39. 
Apron-plates, 140, 153, 154, 
Arch of reverberatories, 96. 
Argall roaster, 73, 
Argentine copfwr output, 15. 
Ar^ smelter, 71, 88. 
An7X)na copper output, 17. 
„ costH of production, 14. 
„ mines, 8, 17. 
„ ores, 46. 



Arnold, J. 0., 34. 

Arrangement of electrodes, 227-229, 234-237. 

tanks, 234, 235. 
Arsenic in copper, 20, 23-25, 28-33, 41, 43, 44, 
207, 217, 230. 
' „ „ -matte, 37, 207, 215, 217. 

„ electro-refining, 215, 217, 231. 

Arsenides, Roasting of, 66. 
Ash-beds, Copper-bearing, 45. 
Aspinall, 34. 
Assaying, 47. 

Associated Copper Smelters, 7. 
Atacamite, 47. 
Atlantic Mine, 45. 
Atmosphere of the blast furnace, 115, 186. 

„ „ reverberatory furnace, 63, 

Auger-Former, 56. 
Austen, Roberts- 34. 
Austin, L. S., 80, 145, 170, 191. 
Australian copper output, 15. / 

„ mines, 7, 8, 15. 
Austria, copper production, 15. 
Azurite, 47. 


Baggaley, 194, 200, 202, 215. 

Balakala, Cal., 75. 

Baltic Mine, 14. 

Baltimore Company's Smelter, 194. 

Barilla, 45. 

Barnett, M., 229, 235, 243. 

Barrel-shaped converters, 193-197. 

Barring of blast furnaces, 124, 127, 130, 158, 

Bases in slags, 148-150, 176, 180, 181. 
Basic converter practice, 51, 181, 193-202, 
„ linings, 193-197, 200, 202, 204, 207, 212- 

„ silicate slags, 148, 149, 176, 179, 180, 183. 
Bauer, 50. 

Bedding systems, 156. 
Beds of refining furnaces, 219. 

„ reverberatory furnaces (see Hearths). 
Bending tests for copper, 20. 
Bengough, G. D., 28. 
Bessemer, 170, 192, 195. 
Bessemerising of copper mattes, 42, 51, 191, 
„ of low-grade matte, 51, 193. 

„ in pyritic smelting, 172-182. 

Best select copper, 40, 42, 44, 214, 215. 
„ selecting process, 9, 40, 214, 215. 
Bi- silicate slags, 148, 149. 
Bismuth in copper, 23, 30-33, 44, 207, 217, 
„ bessemerising, 217. 

„ electro -refining, 231. 

refining, 217, 230. 
Black copper, 43, 44. 

i Black copper smelting, 115, 116. 
' Blast for blast furnaces, 125, 160, 161. 
! „ converters, 195-198, 207, 213, 215. 

! „ pyritic smelting, 178-184, 188. 

1 „ sintering, 58. 

I Blast furnaces, 52, 55. 
I „ furnace, early forms, 10. 
practice, 113-191. 
smelting, 10, 53, 54, 67, 108^ 
110, 111. 
„ „ water-jacketing, 10 (see also 

Water- jacketing). 
„ mains, 141. 
„ pressure, 135, 141, 159, 160, 167, 184, 

„ roasting, 51, 55-59, 80, 110-112, 169. 
Blister copper, 41, 43, 44, 205, 213. 
Blount, 34. 

Blowers for blast furnaces, 141. 
Blowing to blister, 213-215. 

„ fine concentrates into converters, 55. 
of converters, 192, 195, 205-216. 
Boiler tubes, Copper for, 19, 33. 
Boiling in furnace refining, 221. 

„ point, 24. 
Bolivia, copper ores, 46. 

„ „ output, 15. 

Bornite, 46. 
Bosh-angle, 139. 
Boshes, 123, 138. 

Boston and Montana Smelter, 191. 
Bottoms, Copper, 40. 
Bottom-plate for blast furnaces, 136. 
Bottom of reverberatories (see Hearths). 
Bracing of reverberatories, 96, 99, 221. 
Brasque hearths, 117, 123, 134. 
Brasses, 21, 40. 

Breadth of reverberatories, 89. 
Breakdown of electrodes, 230, 233, 237, 238. 
Breast plate, 139, 159, 160. 
Brick furnaces, 124. 
Bridging of blast furnaces, 125-127. 
Brinell test, 32. 
Bringing to pitch, 19, 26, 28, 39, 40, 43, 217, 

222, 223, 230, 238, 241, 242. 
Briquettes in blast-furnace charges, 150-152. 

Coke in, 57, 101. 
Briquetting, 51, 54-57, 169. 
British copper mining, 2, 56. 

„ production, 15. 

Brittleness of crude copper, 242. 
Bronzes, 21. 
Brown, W., 17. 

„ roaster, 70, 71. 
Bruckner roaster, 72, 75. 

,, ,, Introduction of, 10. 

Brunton sampler, 49, 50. 
Bullion in electro-refining (see Values). 
Bustle pipes, 141. 
Butte mining, 8. 
„ ores, 46, 63. 

„ „ Preliminary treatment, 52. 
„ „ Roasting of, 63. 




Cadmium in copper, 23. 

Calculation of blast-furnace charges, 146, 147, 

California, copper output, 17. 

Practice in, 75, 168. 
Calumet and Hecla, 8, 14, 45. 

„ „ costs of production, 14. 

Canada, copper ores, 46. 

„ „ output, 15. 

Cananea, 80. 

blast furnaces, 136, 142, 145, 156, 
„ Costs of production at, 14. 
„ „ roasting, 78. 

„ reverberate ries, 91, 105-108. 
settlers, 163. 
Capacity of blast furnaces, 114, 123-125, 129, 
131, 135, 151, 153, 174, 179, 
converters, 193, 195, 198, 201, 

202, 204, 207, 212. 
electrolytic tanks, 231, 234, 237. 
„ refining"^ furnaces, 219, 224-227, 

reverberatories, 88-92, 99, 102, 
„ roaster furnaces, 70-72, 76-79. 

Cape Copper Company, 68, 73. 
Carbon and copper, 28, 42. 

„ „ silicates, 39. 

,, dioxide and copper, 23. 

„ „ and acid manufacture, 181, 

189, 190. 
„ „ in blast-furnace gases, 181, 

189, 190. 
„ monoxide in copper, 25. 
Carbonaceous fuel in blast furnaces, 115, 119, 

121 (see also Coke). 
Carbonate ores, 47. 
Cars for charging, 141, 153-155, 186. 
Cast copper, 32, 35. 
Casting furnaces, 219, 220, 224, 225. 
machines, 224, 226, 242. 
of copper anodes, 217, 219, 224. 
of merchant copper, 217, 237-242. 
Cathode copper, 39, 43, 44, 230, 233, 238, 
plates, 227-231, 237. 
.sheets, 227, 230-232, 235, 237. 
Cayple-ss, 37, 50. 
Cement copper, 43, 44, 236. 
Cerro de Pa.sco, 75. 
Chalcocite, 46. 
Chalcopyrite, 45-47. 

„ in pyritic smelting, 171. 

„ Roasting of, 65. 

Chambers' briquette machine, 56, 57. 
Changes during besiicmerising, 206-210, 216, 

Channelling in reverberatory grates, 86, 101. 
Channing, J. Parke, 170, 186, 191. 

1 Charge for blast furnaces, 130, 146, 147, 150, 
I 151, 187, 211. 

,, blast roasting, 58, HI, 112. 
„ converters, 212, 213. 
„ „ pvritic smelting, 178-180, 187, 

„ „ reverberatories, 102, 110-112. 
„ „ roaster furnaces, 76-79, 111, 112. 
„ „ smtering, 58, 111, 112. 
„ -cars, 141, 153-156, 186. 
„ -sheets, 153, 157, 158, 187. 
„ trains for blast furnaces, 156-158. 
„ „ reverberatories, 99. 

Charging of blast furnaces, 140, 146, 153-158, 
179, 186, 188. 
of converters, 198, 199, 207, 208, 
„ of refining furnaces, 240. 
„ of reverberatories, 85-87, 91-105 
„ -doors, 141. 
„ -platform, 141. 
Checking of operations, 47. 
Chemical properties, 18, 33. 
Chili bar, 43, 44. 
„ copper ores, 46. 
„ „ production, 15. 
„ Supplies of copper from, 7, 15, 45, 46. 
Chilling in blast-furnace hearths, 115, 117. 
converters, 195, 199, 210, 212. 
„ settlers, 117. 
China, 45. 

Chlorine and copper, 33. 
Chromite linings, 162. 
Circulation of electrolyte, 232-236. 
Claying of reverberatories, 95. 
Clean slags, 131-133, 148, 149, 188, 211. 
Cleaning of electrolytic tanks, 238. 
Cleaning-out of pyritic furnaces, 178-179. 
Clinkering of reverberatory fire-grates, 84, 85, 

Cloud, T. C, 80. 

Coal for reverberatory furnaces, 89, 90, 101- 
102, 105. 
„ consumption in reverberatory work, 84, 
102, 105. 
Coaling of reverberatories (see Firing). 
Cobalt in copper, 23, 217. 
Coinage alloys, 2 1 . 

Coke charging in blast furnaces, 153, 158. 

„ consumption in the blast furnace, 119- 

122, 143, 151, 152, 158, 169, 177, 178, 

182, 187. 

„ in the blast furnace, 121, 129, 143, 147. 

„ in pyritic smelting, 1(J9, 173, 174, 178, 

180-184, 187, 188, 19J. 
„ recovery from reverberatories, 101. 
„ used in briciuettes, 57, 101. 
Cold rolling of copjxir, 31. 
Collection of slimes, 238. 
Colorado, Practice in, 17. 
Colour of converter flame, 206, 207, 200, 210, 
„ copper, 23. 



Colour of matte, 39. 

Commercial copper, 29, 39, 44, 238. 

Composition of anodes, 217, 230-233, 238. 

„ blast-furnace charges, 150- 

„ blast-furnace gas for acid- 

making, 189-191. 
,, blast-furnace slags, 147-153, 

179, 180, 188. 
„ briquettes, 57. 

cathodes, 230, 232, 233. 
„ charges for pyritic smelting, 

173-176, 180, 187. 
„ charges for reverberatory 

smelting, 82, 102, 108- 
„ charges for blast- roasting, 

111, 112. 
„ converter copper, 218. 

copper, 31, 34, 39, 40, 44, 217, 
218, 230, 238. 
„ „ for fireboxes, 19, 20, 

„ mattes, 37, 50, 61-65, 
82, 87, 131, 146, 
147, 175-188, 207, 
„ ,, ores, 45, 46. 

electrolyte, 231-237. 
„ pyritic -smelting slags, 175- 

„ reverberatory charges, 82, 

102, 108-112. 
„ reverberatory matte, 82, 103, 

slag, 103, 104. 
„ roaster products, 79. 

Compounds of copper, 35. 
Concentrates, coarse. Treatment of, 53, 63, 
Fine, 53-55, 61, 63, 68, 108- 
112, 158. 
Concentration in the blast furnace, 57, 113- 
121, 179. 
„ pyritic smelting, 174-183, 

187, 194. 
„ reverberatory smelting, 81, 

smelting, 62, 63, 66, 200. 
Wet, 51, 53, 54, 61, 63. 
Condition of charge for blast roasting. 111, 
„ „ reverberatory smelt- 

ing, 81, 88, 91, 108- 
„ „ roasting, 62, 76, 78, 

„ copper market, 13. 

„ impurities in copper, 26, 31, 34. 

Conditions for electro -refining, 232, 234, 237. 
good settling, 117, 131, 133. 
„ successful acid-making, 189- 


Conditions for successful bessemerising, 192, 
204, 211. 
„ „ pyritic smelting, 

„ „ roasting, 66, 76, 

I Conductivity of electrolyte, 232. 
I „ Electrical," 23, 24. 

I „ High, copper, 18, 22, 39, 41, 44, 

218, 242. 
' „ Thermal, 19, 25, 117, 221, 239. 

! Conglomerate deposits, 45. 
j Conker plate of reverberate ries, 99. 
! Connections in electro -refining, 228, 234. 
t Construction of the blast furnace, 113, 114, 
124, 131, 135. 
„ MacDougal roasters, 73-79. 

„ reverberatories, 87, 96, 97. 

Constituents of blast-furnace charges, 150- 
„ briquettes, 57. 

I „ reverberatory charges, 103, 

j 104, 108-112. 

I „ roaster charges, 76, 79, 109- 

I 112. 

j Constitution of copper matte, 37, 38, 50. 
j Consumption of copper for alloys, 21. 
I „ „ in electrical work, 18. 

; Contacts in electro-refining, 228. 
Contamination of cathodes, 231, 233, 235, 
electrolyte, 219, 230, 233- 
Continuous converting, 181, 202. 

„ working in blast furnaces, 115, 

130, 135. 
„ „ in pyritic smelting, 179, 

„ „ of reverberatories, 85, 87, 

92, 94. 
,, „ of roasters, 72. 

Contraction of area, 20. 
Control in bessemerising, 196, 205, 211-213. 
„ in blast-furnace working, 115, 117, 

131, 178. 
„ in pyritic smelting, 173-183. 
„ in reverberatory smelting, 63, 87, 

,, in roasting, 72, 75. 
„ of matte grade, 82, 175-186, 200, 203, 

„ of operations, 47, 82. 
Converter bars, 42. 

copper, 42-44, 217, 218. 
flames, 205, 207, 209, 210, 216. 
-grade matte, 61, 63, 146, 147, 

-linings, 192, 194, 199-201, 211- 

-matte, 181, 188, 192, 194, 203, 

204, 207, 211, 212, 215-217. 
-practice, 192, 199, 203-216. 
,, -process, 35, 41. 

„ Losses in, 116. 



€onvert€r-slags, 192, 203-205, 211, 212. 

Converters, 192-216. 

Converting, 54, 170, 192-216. 

Cooling of hearth in reverberatories, 83-87, 

91, 96. 
Copper as electrical conductor, 14, 34, 242. 
„ for sta,ys, 19, 41. 
„ for steam pipes, 19. 

glance, 46. 
., High-conductivity, 18, 39. 
„ in blast-furnace charges, 150-152. 
„ in converter slags, 211, 213. 
„ in matte, 147. 
„ in refinery slags, 241. 
„ industry. Present position of, 13. 
„ losses in slags, 115-117, 131, 1.32, 145, 

149, 179-181, 186, 188. 
„ matt«, 36, 37 (see Matte, also Grade of 

„ oxide in copper, 26-29. 
„ oxide in furnace refining, 221, 239- 

„ oxides, 35, 43 (see Oxides). 
„ Properties of, 22-34. 
„ pyrites (see Chakopyrite). 
sand, 45. 
silicates, 35, 39. 

sulphate in electrolyte, 231, 236. 
sulphides, 35, 36. 

„ Roasting of, 64. 
., Uses of, 18-22. 
., Varieties of, 24. 
•Copperhill Smelter, Tennessee, 140, 162, 170, 

184, 186, 188, 190, 214. 
-Cornish copper mining, 5. 

„ ores, 46. 

Corrosion of copper, 33. 

„ reverberatory linings, 95. 
Costs in blast-furnace smelting, 124, 130, 143, 
153, 177, 179. 
„ blast roasting, 58. 

electro-refining, 228, 229, 233. 
oil-fired reverberatories, 81, 107, 

production of copper, 11, 14, 
pyritic smelting, 179, 182, 185. 
reverberatory smelting, 102, 
roasting, 62, 75, 78, 177. 
'Cc>ttrell fume settler, 191. 
Cowpcr-Coles process, 19, 238, 243. 
Cranes, 214, 236, 237. 
Crucible- jackets, 137. 
Crude copper, 43. 
■Crusting in blast-furnace work, 129, 130, 153, 

154, 1.58, 178, 185, 188. 
Cuba, Copper output of, 15. 
Cuprite, 46. 
Cuprous oxide in copper, 26-29, 42. 

„ Properties of, 35. 

Current-density, 229, 232, 233, 237. 

in electro-refining, 227, 229, 232-235. 
leakage, 228. 
Cutter for sampling, 49. 

D.wiD, 193, 214. 

Dean, U. 

Decline of British mining, 9. 

De Lamar refinery, 243. 

Density of copper, 24. 

„ mattes, 39, 179. 
electrolyte, 231. 
slags, 147, 148, 149. 
Depositing tanks (see Tanks). 
Deposition of copper. Electrolytic, 232, 233, 
236, 237. 
„ „ from solution, 33, 43. 

„ of moss copper, 38. 
Destruction of converter linings, 199-204. 
Details of blast furnaces, 113, 131. 
„ MacDougal furnaces, 76. 
,, pyritic smelting furnaces, 184, 185, 
refining furnaces, 220, 224, 239. 
„ reverberatory furnaces, 96-99. 
Development of bessemerising, 192. 

„ blast furnaces, 122-127, 145. 

„ copper industry, 2. 

„ reverberatory furnaces, 89, 

„ roasting furnaces, 69. 

„ ,, practice, 69. 

„ smelting practice, 9, 51. 

Difficulties in black copper smelting, 115, 117. 
in casting, 224. ^ 

in converting, 192, 193. 
in electro-refining, 219. 
in pyritic smelting, 177-181. 
in series-system, 228. 
with gaseous fuel, 108. 
with matte-anodes, 26. 
Diffusion of arsenic in copper, 29. 
Dilution of blast-furnace gases, 189. 
Dimensions of blast furnaces, 123-127, 135, 
184, 185, 188. 
converters, 193, 195, 197, 198, 

202, 204. 
electrodes, 229, 230. 
„ heaps, 68, 

„ refining furnaces, 220-224, 

reverberatories, 84, 88-89, 106. 
„ roasting furnaces, 69, 72, 74, 70. 

tanks, 234. 
Dimorphic change in sulphides, 3(5. 
Disadvantages of basic linings, 214. 

„ bessemerising low-grade 

mattes, 203, 
„ Bruckner roaster, 72. 

„ forced draught, 85, 

„ high current density, 233, 

„ impure anodes, 219, 

roasting, 118, 
„ HcricH-Hystem, 228, 

Hmall fire-graUiH, 85, 
Difloharge of blast furnaces, 138 (bc*o With' 



Discharge of roasters, 72, 78. 

Disposal of blast-furnace products, 146, 159 

(see also Withdrawal). 
Dissolved gases in copper, 25, 26, 31, 40, 42. 
Distribution of blast-furnace charges, 153 156, 
179, 186, 188. 
„ electrolyte, 235. 

„ impurities in copper, 31. 

Direct process, 35. 
Doctoring of furnaces, 178. 
Doors, 141. 
" Dope," 210. 

" Doping " in converting, 210. 
" Doubling " in converting, 212. 
Douglas, James, 215. 

Draft in reverberatory work, 83-87, 93, 94. 
Draft-pressures, 85, 87, 92-94, 101, 102. 
Dry copper, 24, 28, 42, 222, 239, 241, 242. 
„ Characteristics of, 24, 41, 42. 

Ducktown Smelter, Tennessee, 154, 170, 177, 

184, 186, 188, 190. 
DuctiHty of copper, 20, 22, 31, 33. 
Dust, 159, 167, 168, 169, 186, 221. 

„ chambers for blast furnaces, 140, 159, 
167, 168, 190. 

„ „ roasters, 72, 78. 

„ losses (see Flue-dust). 
Dwight- Lloyd sintering machine, 59, 60, 109, 


Earth-oxides in slags, 149, 176, 177. 
Economic factors in acid-making, 189, 190. 
Economy of pyritic process, 177, 181. 
Effects of impurities in copper, 18, 22, 23, 34. 

,, mechanical treatment, 22. 
Efficiency of blast furnaces, 114, 129. 
„ in electro-refining, 229, 234. 

„ of reverberatory furnaces, 114. 

Elastic limit, 33. 
Elasticity in blast-furnace smelting, 114, 115, 

124, 130. 
Electrical conditions (see Conditions). 
„ conductivity, 23, 34. 
„ resistance, 24. 
„ uses, 18. 
Electrodes, 227, 229, 234. 
Electrolyte, 229-231, 235. 
Electrolytic copper, 33, 39, 40, 42, 44, 50, 218. 
„ „ Tough pitch, 41. 

refining, 8, 43, 50, 215-218, 227- 
Elimination of constituents in converting,205, 
207, 210, 215, 217, 222. 
„ impurities from copper-mattes 

50, 215, 217. 
,, impurities in furnace refining, 

221-223, 239. 
., iron in the blast furnace, 118, 

„ sulphur in the blast furnace, 

113, 114, 117, 118, 120, 174. 

Elimination of sulphur in reverberatory 

smelting, 81, 82. 
Elkington, 227. 

Elmore depositing process, 19, 238. 
Elongation, 32. 
Emmons, W. H., 140, 145. 
End-feeding of blast furnaces, 154. 
Engineering and Mining Journal, 17. 

„ progress and copper mining, 6. 

uses of copper, 18, 21, 30, 33, 41^ 
Equilibrium diagrams, 22. 
Erection of blast furnaces, 114. 
Erubescite, 46. 

Evans' Klepetko roaster, 74-79. 
Excrescences on anode-plates, 228, 230, 234,. 

Expansion-openings in reverberatories, 99. 
Expense of roasting, 62, 75, 78, 79, 177. 
Extension of blast furnaces, 124, 131. 
External settling, 113, 117, 122, 123, 131, 133. 
Extraction of values in converting, 194, 200,. 

201, 213-215. 

Factors in slag calculations, 147. 

Fahl-ore, 46. 

raiding, F. J., 191. 

Features of blast-furnace practice, 113, 114. 

Fettling of reverberatories, 95. 

Fine concentrates in the blast furnace, 120. 

Fines, Agglomeration of, 47, 109. 

„ in the blast furnace, 53, 55, 108, 110,. 

114, 120, 140, 154, 158, 167. 
„ in the reverberatory, 81, 108, 109. 
„ Preliminary treatment of, 53-57. 
,, produced on wet-dressing, 52-55. 
„ Roasting of, 66-68, 110. 
Fining of crude copper, 238. 
Fire-boxes (see Grates). 
Fire-box plates, 33, 41, 44. 
Fire-brick linings for settlers, 162. 
Firing of reverberatory furnaces, 83-87, 93,. 
100, 101. 
„ roaster furnaces, 72. 
Flame in converting, 205-210. 

„ furnace refining, 223, 224. 

„ reverberatory furnaces, 86, 89, 91,. 

101, 103, 105-108. 
,, roasting, 69. 
Flanging test for copper, 20. 
Flinn, F. B., 243. 

Fluctuations in price of copper, 11, 12. 
Fluidity of slags, 149, 150, 176. 
Flue- dust, 159, 167-169, 186, 221. 

,, losses in the blast furnace, 55, 159. 
„ ,, roasting 72, 78. 

smelting, 81, 110. 
treatment, 82, 106, 108, 110, Illy. 
140, 167-169. 



Flues of blast furnaces, 140, 159, 168, 186. 

,, reverberatories, 85. 
Fluxes in blast-furnace smelting. 147, 149,211. 
converting, 193, 199, 200, 203, 204, 

pyritic smelting, 173-176, 187. 
,, reverberatory smelting, 76, 88, 110. 

roasting, 79, 88, 110. 
,, smelting, 52. 
Focus of pyritic furnace, 171. 
Forced draught in reverberatories, 85. 
Fore-hearths (see Settlers). 
Formation-temperature of slags, 148, 172, 

175, 176, 180, 181, 183. 
Foimdations for blast furnaces, 135. 

„ refining furnaces, 239. 

,, reverberatories, 96. 

Fracture of copper, 23, 242. 
Freeland, W. H., 170, 186, 188, 191. 

,, charger, 154-158. 
Freezing-point curve for mattes, 38. 
Friedrich, 28. 

Fuel consumption in blast furnaces, 115, 120, 
121, 125, 129, 153, 177, 178. 
., economy in blast furnaces, 114, 115, 129, 
177, 178. 
in blast-furnace practice (see Coke). 
in reverberatory smelting, 81, 83, 84, 

86, 91, 103, 105. 
in roasting, 69, 72, 75, 78, 79. 
value of charges in blast furnaces, 115, 

119, 120, 152, 170, 177, 178, 184. 
value of fine concentrates, 58, 111. 
value of sulphides, 119-122, 152, 170, 
173, 177, 178, 203, 212, 213. 
,, value, Loss of, in roasting, 72, 78. 
Fulton, 50. 

Fume, 67, 140, 158, 159, 168, 191, 216. 
Functions of the blast furnace, 113-116, 131, 
133, 134. 
,. coke in pyritic work, 182. 

,, the converter lining, 193, 199. 

oxygen in copper, 20, 23, 26-33, 
40, 41, 42, 44, 221-223, 238- 
the refining furnace, 221, 223, 

the reverberatory, 81, 82-88. 
„ the roasting furnace, 88. 

Furnaces for refining, 219, 239. 

,, roasting, 69. 

Furnace-refined (tough pitch) copi>er, 41. 
Fusibility of slags, 148-150, 176. 

Ganoue, 43. 

Garfield Smelter, Utah, blast furnaces, 156. 

„ „ converters, 194. 

„ „ reverberatories, 91. 

„ „ roasters, 75, 78. 

Gases dissolved in copper, 25, 26, 31, 40, 42. 

Gases for acid manufacture, 140, 146, 167^ 
181, 186, 188-191. 
in blast-furnace work, 140, 146, 156,. 

in copper, 43. 
in furnace refining, 223. 
in pyritic process, 172, 181, 186. 
in reverberatory smelting, 89-91. 
in roasting, 63, 64, 68, 78, 177. 
Gaseous fuel for reverberatories, 108. 

,, products of the blast furnace, 159,. 
166-168, 181, 182, 188-191. 
German copper mining, 5. 

„ „ production, 15. 

„ ,, smelting, 3. 

„ silver, 21. 
Gibb, A., 37, 50. 
Giroux blast-heater, 144. 
Gobbing of furnaces, 174, 178, 188. 
Gold and copper, 21, 23, 34, 40. 
„ in copper, 44, 217, 218, 226. 

ores, 46, 217. 
„ „ -smelting, 217. 

in electro-refining, 229, 230-233, 238. 

! Goodner, 50 
i Gossan deposits, 45. 
i Gowland, W., 3, 17, 34, 90, 126, 145. 
i Grade of matte, 61-65, 82, 87, 131, 175-194,. 
Control of, 82, 175-186, 200,. 
203, 204. 
Granby smelter, 154, 191. 
Granulation of slags, 165. 
Grate-area in reverberatories, 83-85, 88, 89. 
Grating in reverberatory smelting, 83, 85, 86,. 

92, 101. 
Great Falls Smelter, 108, 191, 195. 
Greenwalt, W. E., 243. 
Green-ore matte, 179, 181, 187, 188. 

smelting, 179, 181, 187, 188. 
Guess, (J. A., 171. 


Haas converter, 215, 216, 

Hallowell, .W. 

Hammering and mechanical proixTtics. 'M. 

Hampe, E., 26, .34. 

Hand calciner, 69. 

Hand-charging of blast furnaces, 153. 

Hand-ladling, 224. 

Handling of electrodes, 230. 

Hardening of copi>cr by hammering. 1, 31. 

„ „ impuriticH. 21. 

Hardness of wjpper, 32. 
Harrington, 37, 50. 
Harvard, F. T., 191. 
Haydcn, 50. 
Heap- roasting, 67. 

Hearth-area of blast furnaces, 127, 129, 130^ 



Hearth of blast furnaces, 123, 127, 129, 134, 
,, of refining furnace, 239. 
,, of reverberatory furnace, 86, 89, 91, 

„ of roasting-furnace, 69. 
Heat, Conductivity for, 25. 

,, losses in blast furnaces, 124, 181. 
,, ,, converting, 212, 213. 

,, ,, reverberatories, 83, 86. 

„ ,, roasters, 75. 

„ settling, 133. 
„ production in converting, 194, 198, 

203-205, 210-213. 
„ production in pyritic smelting, 173, 
174-177, 180-184. 
Heaters of copper, 21. 
Heating air for blast furnaces, 143-145, 181- 

Height of blast furnaces, 136, 137, 188. 

„ „ furnace charges, 153, 172, 

184, 188. 
water-jackets, 123, 136-138. 
Henderson process, 7. 
Herrick, R. L., 191. 
Herreshoff roaster, 74-79. 
Heyn, E., 27, 34, 50. 
Heywood, W. A., 194, 205, 215. 
High-conductivity copper, 18, 39. 
Hill, 28. 

Hioms, A., 24, 28, 34. 
History of copper, 1 . 

„ pyritic smelting, 170. 

Hixon, H., 216. 
Hodge's charge-car, 154, 155. 
Hoopfner, 238. 

Hofman, H. O., 37, 50, 80, 243. 
Holwav, 119, 170, 172, 191-193. 
Hood of converter, 197, 198, 201. 
Hoppers in roaster furnaces, 69. 
Hopper- feed to blast furnaces, 140. 
Horse-shoe roasting furnace, 71. 
Hot-blast stoves, 144, 145. 
Hot charges for reverberatories, 87, 92, 
Hughes, G., 34. 
Hungary copper output, 15. 
Hydrocarbons in copper, 25. 
Hydrochloric acid and copper, 33. 
Hydrogen dissolved in copper, 25. 

Idaho copper output, 15. 

Improvements in roasting practice, 69-79. 

Impure anodes, 228-231, 235. 

Impurities and annealing temperature, 31. 

„ and conductivity, 23, 24, 34. 

„ and hardness, 32. 

„ and specific gravity, 25. 

„ elimination from mattes, 50, 217. 

in anode copper, 219, 228-231, 
233, 235. 

Impurities in copper, 18, 20, 23, 24, 26, 28, 
31, 40-42, 217, 218. 
in electro-refining, 219, 228-231, 
233, 236. 
„ in furnace refining, 239, 241. 

Increase in blast-furnace size, 113. 
„ reverberatories, 88-91. 

Incrustation in blast furnaces, 129. 
Industrial uses of copper, 18, 21. 
Influence of conditions on properties of 

copper, 30, 34. 
Ingot copper, 242. 

Intermittent working of roasters, 72. 
Internal settling, 123, 133, 134. 
Iron and copper silicates, 39. 
„ in blast-furnace charges, 118, 147, 149- 

,, in converting practice, 193, 200, 203, 

205, 207, 212-217. 
„ in copper, 20, 23, 31, 44. 
„ in electro-refining, 229-232. 
„ in furnace refining, 221, 241. 
„ in pyritic smelting, 172-176, 179, 180, 

181, 183, 188. 
„ in slags, 149-153, 211. 
,, pyrites (see Pyrites). 
„ reduction in the blast furnace, 115, 116. 
„ silicates, 148-150. 
„ sulphide in matte, 37, 146, 147, 203. 
„ „ reactions on roasting, 64. 

Isabella Smelter (see Ducktown). 
Italy, Copper output of, 15. 

Jacketting (see Water -jacketting). 
James (and Nicholl) process, 9, 35. 
Japan, Copper ores of, 46. 
,, „ output of, 15. 

„ old smelting methods, 3. 
Johnson, F., 26, 34, 243. 
T., 34. 
„ on electrical conductivity, 23, 24. 


Katanga, 40. 

Keith, 238. 

KeUer, E., 50, 71, 193, 202, 215, 217, 243. 
I Keswick Smelter, Cal., 119, 120. 
; Kiddie blast heater, 144. 
I Kilns, Roasting in, 67, 68. 
j Kletko (and Evans) roaster, 74-79. 

Knudsen process, 194, 215. 

Labour for blast furnaces, 130. 

,, refining furnaces, 224. 

,, reverberatories, 100. 

roasters, 69, 70, 72, 75, 78, 79, 


Ladles, 16C. 

„ for refining furnace, 224-22C, 242. 
Ladling of copper, 224. 

Lake copper, 40, 42, 44, 45. ■ 

„ Superior ores, 43. , 

Lamb, S., 24, 34. 

Lambert, 7. I 

Large reverberatories, 81, 87, 88. ; 

Launders for matte, 116. i 

for slag, 164, 165. ; 

,, -castings, 166. 
Law, E. F., 34. j 

Lawrie, 34. ; 

Lay-out of blast-furnace nlant, 156. ' 

Leaching processes, 67.4' 

Lead anodes, 236. ■ 

„ in copper, 20, 23, 30, 32, 44, 217. 

-matte, 37, 215-217. 
„ in blast-furnace fume, 168. 
„ sulphide, Roasting of, 66. 
Leakage of air in blast-furnace blowers, 141, 
„ of air in reverberatory furnaces, 83, 

85, 86. 
„ in electro-refining, 228. 
„ in water-jackets, 139, 158. 
Length of blast furnaces, 131, 135. 

„ reverberatories, 89, 91. 
Lengthening of blast furnaces, 124-127, 131. 
Levelling of charges in reverberatory smelting, 

Levy, D. M., 216. 
Liberating tanks, 236. 

Lime in blast-furnace charges, 151-153, 180. 
,, blast-roasting process, 58. 

pyritic process, 180, 181, 186-188. 
., roaster charges, 76, 77, 88. 
slags, 149-151, 153, 180. 
Linings for converters, 192-194, 197-201, 211- 

Lining of converters, 201-204, 
settlers, 162, 163. 
tanks, 228, 229, 234. 
Liquation of sulphides, 172, 182, 194. 
Liquid fuel for reverberatories (see Oil). 
Lloyd-Dwight sintering machine, 59. 
Loam lining for settlers, 162, 163. 
Locomotive work. Copper for, 20. 
Losses by fines in the blast furnace, o5, 110. 
„ by fines in the reverberatory, 110. 
„ in blast-furnace smelting, 1 15-1 17, 125, 

146, 152, 179. 
., in converting, 195, 200, 204, 209, 210, 

212 214 222 
„ in pyritic smelting, 179-181, 186, 188. 
„ in refining, 116, 221, 222. 
„ in roasting, 62, 67, 68, 177. 
„ of heat (j»ee Heat losses). 
LouiH, H. (and Schnabel), 243. 
Low-grade mattes, 181, 200, 203,204,212-214. 

„ ores, 52. 
Lugs of electrodes, 226, 229, 230, 236, 
Lump ores. Roasting of, 66, 67. 


M.vcDouG.\L roasters, 73-79. 

M'Murty- Rogers sintering process, 58, 80. 

Magnesia in slags, 150. 

Magnesite linings (see Basic linings). 

Malachite, 46. 

MaUeability, 1, 22, 31, 33. 

Management of converters, 212. 

Manganese and copper, 21. 23. 

Manlies, 193, 200. 

Mansfeld furnace, 122. 

Mantle plates, 140, 153, 154. 

Marchese, 238. 

Mass inrtuence, 194. 

Mathewson, E. P., 89, 125, 127, 145, 207, 211„ 

Matte, 36, 37, 50, 61, 118 (see also Grade of 
,. anodes, 238. 
„ for converters, 134, 135, 192, 200, 202- 

„ in pyritic smelting, 178, 181, 183, 186,. 

„ -pool in reverberatory smelting, 87^ 

91, 94-97. 
„ Propertiesof, 38, 39, 91. 
Mechanical charging of blast furnaces, 140, 
153, 167, 188. 
,, operation of roasters, 69-75. 

„ ■ properties of cop})er, 18, 30, 33,. 
34, 43, 45, 238, 241, 242. 
rabbling, 69-75. 

treatment, Effects of, 20, 30, 32,. 
Mechanism of casting ladles, 225-226. 

,, furnace refining process, 223. 

„ pyritic process, 171, 172, 183. 

Melacomite, 46. 

Melting agent, Blast furnace as, 113, 114. 
„ function of blast furnace, 114, 117. 
„ point of copper, 24, 117. 
„ „ copper- mattes, 37, 38. 

„ „ slags, 148. 

Merchant copper, 238, 239. 
Metallic copper in blast-furnace slags, 116. 
Metallography, 26. 
Mexico copijer ores, 46. 

,, „ output, 15. 

Michigan copper output, 17, 
MUton, J. T., 34. 
Mineral industry, 17, 191, 243. 

„ statistics, 17. 
Mining in Britain, 2, 5, (>. 
,, Germany, 5. 
„ Spain, 5, 7, 8, 
MitcheU blast heater, 144. 
Mixed silicate slags, 149, 150. 
Mixing of samples, 49. 
Modem smelting practice, 37, 51, 61, 62, 66, 

„ „ „ Development of, 9. 

„ reverberatory smelting, 87. 



Modifications of converter practice, 214, 216. 

„ electro-refining, 238. 

Moisture in briquettes, 58. 
Monel metal, 21. 

Monopolies in copper industry, 11. 
Mono -silicates, 149, 150. 
Montana copper ores, 8, 46, 52, 63. 

„ „ output, 17. 

Moore, R. R., 80, 215. 
Moss copper, 38. 

Moulds of casting furnaces, 224-226. 
Mount Lyell, Tasmania, 171, 172, 174, 195. 
Mud (see Slimes). 
Muffles, 69. 

Multiple system, 227-229. 
Muntz metal, 21. 


Native ores, 43, 45. 

„ preliminary treatment, 51. 

Natural draft in reverberatories, 85, 
Neal, C. B., 215. 
Nevada, 8, 14, 17, 75. 

„ copper output, 17. 
Newfoundland copper output, 15. 
New Mexico copper ores, 45. 

„ „ output, 17. 

NichoU and James process, 7, 35. 
NichoUs, F. S., 191. , » 

Nickel and copper, 34. 

in bessemerising, 205. 

in copper, 20, 23, 33, 44, 217. 
mattes, 37, 205. 

in electro-refining, 230. 

in refining, 217, 230. 
Nitric acid and copper, 33. 
Nitrogen in blast-furnace gases, 189. 
" Normal " converter practice, 211. 
North American supply, 15, 17, 45. 

„ Carolina ores, 46. 
Norway, copper output, 15. 
Nose-pieces of spouts, 159, 160, 164. 
Notches (see Slag notches). 

Objections to external settling, 133. 
,, internal settling, 134. 

„ roasting, 62, 67. 

„ water- jacketing, 124, 

'Objects of electrolytic refining, 218. 

„ furnace refining, 223. 

OfEerhaus, C, 93, 112, 191, 216. 
Off- takes of blast furnaces, 140, 167. 
O'Hara calciner, 70. 
Oil burners, 106, 107. 
„ fuel for reverberatory furnaces, 81, 91, 
105, 106-108. 
Open-air roasting, 67. 

Operation of the blast furnace, 114, 146, 158, 

i Operation of converters, 192, 207, 210-214. 
„ electro-refining process, 227, 

230, 236. 
„ furnace refining process, 221, 

223, 238-243. 
} „ large reverberatories, 81. 

Ore bedding, 156. 
„ of copper, 43-47. 
„ for converter fluxes, 213, 214. 
linings, 201, 214. 
,, for pjritic process, 171, 176, 177, 184, 

187, 188. 
„ -lining for settlers, 162, 163, 186. 
Organisation in bessemerising, 214. 

„ at smelters, 47, 75. 

Outlets of blast furnaces, 160. 

settlers, 164-166. 
Output of blast furnaces, 124, 129-131, 133, 
135, 188. 
„ copper, 15-17. 

electrolytic tanks, 228, 229, 232, 
234, 237. 
Over-fire, 176, 184. 

Over-poled copper, 24, 26, 42, 50, 242. 
Oxidation in bessemerising, 194, 195, 200, 205, 
207, 208, 210, 211, 214, 217. 
in the blast furnace, 113-122, 171, 
in electro-refining, 231, 232. 
in furnace refining, 221, 222, 239- 
„ in pyritic smelting, 174-184. 

„ in smelting, 62. 
„ reactions in roasting, 64, 65. 
Oxides in copper, 26-29, 23^^42.4^ 
„ of copper, 35, 43. ^ 
„ ores, 43, 45-47. ^' 
„ „ Preliminary treatment of, 51.^ 
Oxidised constituents of the blast-furnace 

charge, 113-115, 118, 119. 
Oxland roaster, 72, 73. 
Oxygen and copper, 33, 40, 41. 

„ in copper, 20, 23, 26-33, 42, 44, 218, 

„ in furnace gases, 171, 172, 182, 189, 
in furnace refining, 221, 222, 239- 
„ in gases for acid manufacture, 189- 
in pyritic smelting, 172-175, 182, 

ratio, 147, 150. 

Panels of water-jackets, 129, 137, 185. 

Parallel system, 227-229. 

Parkes' roaster, 73. 

Parrott Smelter, Butte, 125, 193. 

Partial pyritic smelting, 121, 143, 147, 149, 

158, 169, 177, 184. 
Peacock ore, 46. 




Pearce, R., 88, 92. 

,, roasting furnace, 71. 
Percy, John, 2, 17, 34. 
Perth Amboy Refinery, N.J., 194, 234. 
Peru copper output, 15. 

„ smelting practice, 75. 
Peters, E. D., 65, 80, 110, 112, 143, 145, 170- 

172, 183, 191, 220, 243. 
Philp, 37, 50. 

Phosphorus in copper, 23, 242. 
Physical properties of copper, 18. 
Pierce (and Smith), 194, 202, 215. 
Piltz, 123. 

,, blast furnace, 10. 
Pipe stoves, 144. 
Pitch of copper, 224. 
., Bringing to, 19, 26, 28, 39, 40, 43, 217, 
222, 223, 238, 239-242. 
Pittsmont Smelter, Butte, 193. 
Plant for acid making, 190, 191. 
Platforms for charging blast furnaces, 141. 

., of reverberatory furnaces, 86, 100. 
Platinum in copper, 23. 
Poling of copper, 25, 42, 222, 223, 239, 240, 

242, 243. 
Polybasic slags, 150, 176, 181. 
' Porphyry " camps, 14. 
Portuguese copper mines, 7, 15. 

„ „ output, 15. 

Pots for blast roasting, 58. 
Power for electro-refining, 233, 234. 
Precipitation of copper from solution, 33, 43. 

impurities, 232, 233. 
Preliminary refining of copper, 217, 219, 222, 
223, 228-230. 
fines, 53, 61, 110-112'. 
ores, 47-49, 51, 54-61. 
Preparation of anodes, 217, 219, 230. 
cathodes, 230. 
„ floor for heap- roasting, 68. 

Pressing of briquettes, 27. 
Pressure of blast for blast furnaces, 135, 141, 

159, 160, 167, 184, 188. 
Prevention of heat losses in reverberatory 
smelting, 86. 
„ losses, 116. 

Price of copper, 11, 40. 
Primitive smelting methods, 3. 
Principles of converting, 194, 215. 
„ copper smelting, 51, 61. 

„ electro-refining, 229. 

„ the MacDougal roaster, 73. 

„ reverberatory smeltiiig, 81, 83, 

„ the Welsh process, 9. 

Production, Statistics of, 5, 15-17. 
Products of the blast furnace, 114, 117, 146, 
159, 186, 214. 
blast roasting, 58, 110-112. 
converting, 193, 199, 205, 209, 
210, 212. 
Properties of copper, 18, 22, 34. 

matte, 38, 39. 

Properties of dry copper, 24, 41, 42, 239, 241- 

„ Mechanical, 30. 
Physical, 18. 
Prosser roaster, 71. 

Pulverised fuel for reverberatories, 105. 
Punching of tuyeres, 142, 158, 195, 198, 206, 

210, 213, 214. 
Purification of copper, 217-243. 

electrolyte, 230, 235, 236. 
Pyrites, 45. 

„ reactions on roasting, 64. 
Pyritic effect, 121, 122. 

„ principle, 62, 113, 118-122, 158, 169. 
„ smelting practice, 67, 121, 140, 143, 
145, 146, 158, 169-188, 191-194, 
Pyrrhottite, 45. 

„ reactions on roasting, 64. 


QuiNCY mine, 45. 

Rabbles of MacDougal furnaces, 76, 77. 
Rabbling of roaster furnaces, 69-71. 
Rachette furnace, 10, 125. 
Radiation losses in reverberatory smelting, 

83, 86. 
Rapidity of smelting in blast furnaces, 124, 

129, 133, 135, 143, 
159, 160. 
„ ,, reverberatories,83-92. 

Rate of deposition, 232. 
Reactions in the blast furnace, 113, 117-119. 
converting, 194, 205, 207, 210. 
,, furnace refining 239-243. 

„ pyritic smelting, 171, 172, 178, 

180, 182, 183. 
,, reverberatory smelting, 81, 114, 

roasting, 36, 63, 64, 66, 109-1 12. 
,, sintering, 57, 110-111. 

smelting, 61, 62. 
Re-concentration in pyritic smelting, 179, 

181, 186-188. 
Recovery of copper from slags and residues, 

116, 119,211. 
Rectangular blast furnaces, 123. 
Reducing gases in annealing, 28. 
Reduction in the blast furnace, 113, 115-117. 
„ of oxides in furnace refining, 223- 
smelting, 113, 116, 117, 120, 122. 
; Rcdnithite, 46. 

References, Lists of, 17, 34, 50, 80, 112, 191, 
I 215, 243. 
Refinery slags, 221, 222. 
Refining of copper, 25, 26, 00, 54, 211, 214, 



Refining, Electrolytic, 8, 43, 50, 215-218, 227- 
„ of electrolytic copper, 50, 238-343. 
„ furnaces, 219, 220, 224, 239. 
Losses in, 116, 221, 222. 
Regeneration of electroljrte, 236. 
Regulation of reverberatory furnace working, 

87, 92. 
Re-lining of converters, 199, 200, 204, 213, 

Removal of blast-furnace products (see With- 
„ impurities in converting, 207, 215. 

„ „ furnace refining, 221, 

Renwick, C. W., 191. 
Repairs in roaster furnaces, 79. 

,, reverberatory furnaces, 95. 
Replacing jackets of blast furnaces, 127, 158. 

„ values of bases, 150, 180. 
Requirements for good blast-furnace slags, 
,, good reverberatory prac- 

tice, 81, 87, 109-112. 
„ refining furnaces, 221. 

,, roasting furnaces, 69. 

„ successful pyritic practice, 

Resistance, Electrical, 24. 

of electrolyte, 232, 233. 
Reverberatory fore-hearths, 135. 

furnaces, 52, 69, 88. 
smelting, 54, 00, 61, 63, 76, 80- 
112, 215. 
„ compared with blast 
furnace, 114, 117. 
„ efficiency, 114, 124. 
„ reactions, 81, 114, 
119, 215. 
Rice, C. T., 191. 
Richards, 53. 

Rickard, T. A., 145, 170, 191. 
Ricketts, L. D., 78, 80, 105-108, 112. 
Rigidity of arsenical copper, 33, 41. 
Rio Tinto, 8, 67. 
Roaster gases, 63. 

,, process, 35. 
Roasting, 36, 47, 51, 54, 55, 61-68, 88, 109- 
112, 177. 
in heaps, 67. 

in pyritic process, 176, 182. 
Objections to, 62, 67. 
Open-air, 67. 
practice, 66-71, 82, 87, 88. 

„ early improvements, 10. 
preliminary to blast-furnace treat- 
ment, 113, 115, 118. 
reactions in sintering, 57, 110, 111. 

on, 36, 63-65, 109-112. 
yards, 67. 
Roberts- Austen, Sir W. C, 34. 
Rogers (and M'Murty) sintering process, 58. 
Rontgen, 37, 50. 

Roofs of reverberatories, 99. 

Ropp roaster, 71. 

Rotary blowers, 141. 

Rotating furnaces for roasting. 72, 73 

Rudeloff, E., 34. 

Russia, Copper output of, 15. 

Sackett, B. L, 191. 
Sample-cutter, 49. 
Samphng, 47, 48, 50, 54. 
Costs of, 79. 
„ from blast furnaces, 166. 
„ from converters, 213, 243. 

in furnace refining, 222, 223, 226,. 
227, 239, 241, 242. 
Sand, Copper, 45. 
Scrap in electro-refining, 228, 237. 
Schnabel, C. (and Louis), 243. 
Schreyer, F., 215. 
Seasoning of basic converters, 203. 
„ refining furnaces, 240. 

Secretan combination, 1 1 . 
Sectioning of blast furnaces, 123, 124, 137,. 

158, 185. 
Selenium in copper, 26, 30, 207, 217. 

,, electro-refining, 232. 
Semenikow, 192. 
Semple, C. C, 216. 

Separation of matte and slag (see Settling). 
Series-system, 227-229. 
Sesqui-silicates, 149. 
Settlers, 127, 130, 160, 162-166. 
Settling, 39, 113, 116, 117, 122, 123, 127, 128- 
135, 147, 158, 179, 181, 186, 211, 
„ (in reverberatory furnaces), 81, 91, 
Shaft of blast furnaces, 122. 
" Shaking-out " of gases, 241. 
Shape of blast furnace, 123, 125, 135. 
Shelby, G., 145, 150, 191. 
I ,, blast-furnace top, 140. 
! „ oil burner, 106, 107. 
Short-circuiting in electro-refining, 219, 227, 

228, 233, 234. 
Shots from converters, 205, 210. 
Siemens gas-fired furnace, 108. 
Siemens-Halske process, 238. 
Silica in blast-furnace charges, 150-153, 180, 
187, 188, 191. 
„ in converting, 194, 200-203, 212, 213. 
,, in pyritic process, 172-182. 
„ in slags, 148-153, 180, 181, 191. 
„ -lined converter, 193-197, 200, 201, 205, 
Silicates in the blast furnace, 116, 119, 148- 
152, 180, 181, 187, 188, 191. 
,, of copper, 35, 39. 

148-153, 180, 181, 187, 188, 




Siliceous flux in converting, 194, 200-203, 212, 
linings of converter, 193-197, 200, 
201, 205, 207, 211-213. 
Silicon in copper, 23, 242. 
Silver and copper, 21, 23, 24, 218. 
,, in blast-fumace fume, 168. 
in copper, 44, 217, 226. 

matte, 37, 217. 
ores, 46, 217. 
,, in electro-retining (see Values). 
Sintering, 51, 55, 58, 59, 80, 110, 111, 112, 

Sites for heap-roasting, 67. 
Situation of electro-refineries, 8, 9, 218. 
Size of blast furnaces, 113, 122, 125, 135, 184, 
185, 188. 
„ converters, 193, 195, 197, 198, 202, 

„ electrodes, 229, 239. 
„ grates in reverberatories, 84. 
„ material for blast furnaces, 55, 110. 

„ roasting, 59. 
„ „ roasting, 68. 

„ refining fmnaces, 220, 224. 
„ reverberatories, 84, 87-90, 105. 
„ tanks, 234. 
Skimming of refining furnaces, 240, 241. 

,, reverberatory furnaces, 94, 95, 

Slags, 35, 39, 62. 

„ in blast-fumace smelting, 116, 129-133, 

145, 147-154, 180, 181, 187-191. 
„ in converting, 192, 203-205, 209-213. 
„ in furnace refining, 222, 241. 
in pyritic smelting, 172-188. 
„ in refining, 221. 

,, in reverberatory smelting, 94, 95, 102- 
Slag-formation in roasting, 88. 
,, -foundations for reverberatories, 96. 
„ -notch of blast furnaces, 136, 158-162. 
„ -spouts, 128, 158-166. 
Slagging-stage in bessemerising, 205, 207. 
Slimes in briquetting, 57, 152. 

„ in electro-refining, 228-234, 236, 238. 
„ Treatment of, 53-56, 152. 
Slotted tuyeres, 185. 
Smelting practice, 51, 54, 81, 204. 

„ scheme at Anaconda, 54, 55. 
Smith (and Pierce converter), 194, 202, 215. 
Smoke problem, 171, 188. 
Soluble constituents of electrodes, 229, 231. 
Solubility of copper in iron sulphide, 38. 
„ slags, 116. 
gases in copper, 25, 28. 
matte in slag, 132. 
sulphides in slag, 116, 149, 179. 
Sound anodes, 229, 230. 
Sources of copper, 43. 
South Wales, 4, 5, 71, 73. 
Span of reverberatory arch, 89. 
Spanish copper mining, 7, 15. 


Spanish copper ores, 46. 

„ production, 15. 

Special bronzes, 21. 
Specific gravity of copper, 24. 

„ mattes, 39, 
„ electrolyte, 231 . 

slags, 147-149. 
Specifications for copper, 19, 20. 

M „ for firebox plates, 

„ „ for Post Office work, 

Speculation in copper markets, 11, 12. 
" Spewing " of copper, 222. 
Spindles of MacDougal roasters, 74, 79. 
Spouts, 128, 158-166. 
Sprouting of copper, 222. 
Stacks of blast furnaces, 140, 169. 

„ reverberatories, 89. 
Staffordshire, Copper smelting in, 4. 
Stages in converting, 205, 209, 210, 213, 216. 

refining, 223. 
Stahl, 50. 

Stamping of briquettes, 57. 
Stamp-milling of native copper ores, 51. 
Statistics of copper, 15. 
Staying of reverberatories, 96, 99, 221. 
Stays, Copper for, 19, 41. 
Steam-pipes, Copper for, 19. 
Steptoe Smelter, Nevada, 75. 
Stevens, H. J., 17. 
Sticht, R., 171, 172, 182, 183, 191. 
Storing of matte in blast furnaces. 134. 
„ „ reverberatories, 94. 

„ „ settlers, 135. 

Stoves for heating blast, 145. 
Straightening of anodes, 228. 
Strength of cathode copper, 18, 232, 238, 
„ copper, 21, 32, 33, 41, 238, 239, 

„ electrolyte solutions, 231. 

Stripping of cathodes, 228, 230. 
Sub-silicates, 147. 
Success in pyritic smelting, 181. 
Suliftelma Smelter, Norway, 194. 
Sulphide ores, 43, 45. 

„ Preliminary treatment of, 52. 

Sulphides, Fuel values of, 119-122, 152, 170, 
173, 177, 178, 203, 212, 213. 
„ in the blast furnace, 113, 115, 117- 
121, 147, 158, 171-176, 180, 
189, 191. 
„ of copper, 35-37, 39, 146, 147 
Sulphur and copper, 33, 36, ft], 146, 147, 
„ dioxide in blast-fumace gases, 188- 
copper, 25, 43, 222. 
„ furnace gases, 189, 190. 
furnace refining, 221-223. 
„ „ gases for acid-making. 1K9, 







Sulphur dioxide in pyritic smelting practice, 
182. * 

„ elimination in blast-fumace smelting, 
113, 114, 117, 118, 120, 146, 147, 
elimination in blast roasting, 58. 
elimination in reverberatory smelt- 
ing, 81. 
elimination in roasting, 65, 77, 78, 

88, 110, 177. 
in blast-fumace charges, 146, 147, 

150-152, 177, 189. 
in blast-furnace smelting, 116, 120, 

146, 147, 189. 
in converting, 194, 205, 207, 211, 

212 215 217 
in copper, 20, 23, 31, 32, 44, 222. 
in electro-refining, 230. 
in furnace refining, 221. 
in pyritic smelting, 171-173, 177, 

in roasted products, 65, 70, 102, 104, 
Sulphuric acid and copper, 33. 

in electrolyte, 231, 236. 
„ manufacture, 146, 147, 181, 

Superior (Lake) copper, 40. 

mining, 7, 8. 
„ „ ores, 43, 45. 

„ „ production, 15. 

,, costs of production, 14. 

„ treatment of ores, 51. 

Superstructure of blast furnaces, 140, 168, 

Suspension of electrodes, 230, 231, 234-237. 
Swansea smelters, 4, 5, 71, 73. 
Sweden, Copper output of, 15. 
Systems of working in converting, 192, 211, 
„ „ electro -refining, 227, 


Tamarack Mine, 45. 
Tanganyika ores, 8, 45. 
Tank efficiency, 229. 

„ liquors, 219, 230, 231. 
„ slimes, 219, 231, 232. 
Tanks for electro-refining, 219, 227-230, 234, 

Tasmanian mining, 8. 

Tap-holes of blastfurnaces, 127, 136, 159-161. 
„ refining furnaces, 224, 225. 

„ reverberatories, 101, 103. 

settlers, 152-166. 
Tapping of blast furnaces, 122, 135, 159, 160, 
165, 166. 
of refining furnaces, 224, 225, 242. 
„ of reverberatories, 94, 95, 100, 101, 

Tapping of settlers, 166. 

„ -breast of blast furnaces, 136, 159, 

„ -piece of settlers, 165, 166. 
,, -plate of reverberatories, 103. 

settlers, 165, 166. 
„ -slot of refining furnaces, 224, 225, 
Telegraph, Use of copper for, 19. 
Telephone, Use of copper for, 19. 
Tellurimn in copper, 23, 26, 30-32, 207, 217. 

„ electro -refining, 232. 

Temperature, Effects of, on strength, 20, 22, 
for annealing, 31. 
for roasting, 69, 70, 77, 78. 
for reverberatory smelting, 83, 
84, 87, 89, 91, 94, 96, 99-105. 
for settling, 117, 123, 130, 133. 
in blast-fumace smelting, 123, 

129, 133, 136, 143, 188. 
in converting, 194, 199, 203, 

205, 207, 210-214. 
in electro-refining, 231-233, 235, 

in furnace refining, 224, 225, 

in pyritic smelting, 172-176, 
179, 180, 183, 188. 
Tenacity of copper, 20, 32, 33, 241. 
Tennessee, acid-making, 189-191. 
,, blast furnaces, 140. 

„ converting practice, 202, 211, 214. 

„ Costs of production at, 14. 

„ mining, 8. 

,, ores, 46. 

„ ,, Treatment of, 52. 

„ pyritic smelting practice, 171, 174, 

179, 184-188, 191. 
Roasting at, 62, 68. 
settlers, 162, 164, 165. 
Tensile strength, 20, 32, 33, 241. 
Testing of refined copper, 223, 241, 242. 
Tests during bessemerising, 210, 213. 

„ for copper, 20. 
Tetrahedrite, 46. 
Textile work, Copper in, 21. 
Thermal conductivity, 19, 22, 25. 
Tiers of water-jackets, 137. 
Tilt Cove, Newfoundland, 171, 191. 
Timber for poling, 222, 223. 
Time-element, 194. 
Tin-copper alloys, 21, 34. 
Tin in copper, 20-23, 32, 33, 44. 
Tomlinson, 34. 

Tooele, Utah, Practice at, 75, 91, 211, 214. 
Tops of blast furnaces, 140, 144, 145, 186. 

„ charge in blast furnace, 158. 
Toston, Montana, 170. 
Tough (tough-pitch) copper, 24, 31, 33, 40, 41, 

42,44,223, 231, 238, 239. 
Toughening of copper, 238-241. 
Toughness of copper, 22, 25, 40, 241. 
Trapping of blast, 159, 160. 




Treatment of converter slags, 211. 
„ copper ores, 47, 50. 

„ tine concentrates, 110-112. 

flue-dust, 82, 106, 108, 110, 111. 

slimes in electro -refining, 238. 
True pyritic smelting, 121, 122, 143, 146, 147. 

173, 176, 178-184, 189. 
Turkey, Copper output of, 15. 
TurnbuU (and W. Brown), 17. 
Tuyeres, 122, 123, 141, 142, 158, 185, 188. 

for converters, 192, 193, 195, 197, 
198, 215. 
Tuyere holes in water-jackets, 137, 140. 

-jackets of blast furnaces, 138, 140. 
., -pieces, 140. 
,, -zone, 171, 174-185. 
Tyee Smelter, B.C., Hot blast, 144. 
„ ,, Roasting at, 68. 

Typical pyritic smelting charge-sheets, 206. 
„ reverberatory -furnace charge, 82. 


Ulke, T., 229, 237, 243. 

United States copper output, 15-17. 

Uses of copper, 18. 

„ „ alloys, 21. 

Utah, copper mining, 8. 

M „ output, 17. 

„ costs of production, 14. 

„ smelting practice, 75. 
Utilisation of heat in reverberatory work, 83, 


Vail, R. H., 215. 

Values in anode copper, 230, 

„ converter copper, 43, 215, 218. 

slags, 211. 
„ copper ores, 46. 
„ electro-refining, 229-231, 233, 238. 
Varieties of commercial copper, 39, 40. 
Vats for electro-refining, 219, 227-230, 234, 

Vein deposits in Superior, 45. 
Vertical converters, 195. 
Virginia copper ores, 46. 
Viscosity of slags, 148-150. 
Volatile hydrocarbons in reverberatory smelt- 
ing, 85, 86, 105. 
Voltage for electro-refining, 227, 228, 232. 

234, 238. 
Vortex converter, 215, 216. 


VValkeh, a. L., 34, 191, 224, 226. 
Wallaroo sintering process, 58. 
Walls of reverberatory furnaces, 99. 

WanjukoflF, W., 132. 

Washoe Smelter (see Anaconda). 

Waste heat in reverberatory -furnace gases, 83 

87, 107, 144. 
I „ UtUisation of, 83, 87, 107, 143. 

I Water and copper, 33. 

„ in blast roasting, 58. 
I „ -cooling in roaster furnaces, 74-79. 

„ -jacketting of blast furnaces, 10, 113, 
j 122-124, 134-139, 158, 185. 

! „ -supply for jackets, 124, 137, 139, 140. 
Watson, D., 34. 
' Webb, 34. 
Welsh process, 9, 40, 214, 219, 222-224. 

„ smelting, 82. 
Wethey roaster, 71. 

Wet concentration of ores, 47, 50, 53-55, 61, 
„ processes, 43. 
White-HoweU roaster, 73. 
White-metal, 40, 205, 210, 212, 213, 216. 

stage, 209, 211-214. 
White roaster, 73. 
Width of blast furnaces, 125, 136, 184, 186. 

„ reverberatories, 89. 
Wintle, F. H., 191. 

Wire, Preparation of copper, 19, 238, 243. 
„ Strength of copper, 32. 
,, -bar copper, 40, 242. 
Withdrawal of products from blast furnaces, 

124, 130, 133, 
135, 158-171. 
M if reverberatories, 

83, 86, 94. 
Work, Effects of mechanical, 32. 
Working of blast furnaces, 114, 146, 158, 178. 
„ converters, 178, 185, 186, 188. 

„ electro-refining plant, 227, 230, 

„ largo reverberatories, 84, 92, 97, 

„ MacDougal roasters, 76-80. 

„ pyritic process, 178, 185, 186, 188. 

Wraith, W., 243. 
Wright, L. T., 132, 145, 191. 
Wyoming, Copper production of, 17. 

Yunnan, Native copper from, 45. 

ZiNC-coi'PEE alloys, 21, 22, 34. 
-sulphide. Roasting of, 66. 
in copper, 23, 33. 
in eloctro-rofining, 229, 230. 
in mattes, 37, 217. 
in slags, 150. 


PlNDa^sa btwi. auj ^o W( 





Levy, Donald M 

Modern copper smelting