MODERN COPPER SMELTING.

The Colour of the Converter Flame during the
Bessemerising of Copper Matte.

Fig. 1. Flame at commencement of blow.

Fig. 2. Flame during the first or “slagging” stage.

Fig. 3. “White Metal stage.” Slagging of Copper.

Fig. 4. Flame during second or “blowing to blister” stage.

MODERN
COPPER SMELTING.

BEING
LECTURES DELIVERED AT BIRMINGHAM UNIVERSITY
GREATLY EXTENDED AND ADAPTED, AND WITH
AN INTRODUCTION ON THE HISTORY, USES
AND PROPERTIES OF COPPER.

BY
DONALD M. LEVY, M.Sc., Assoc. R.S.M.,
ASSISTANT LECTURER IN METALLURGY,
UNIVERSITY OF BIRMINGHAM.

With frontispiece, and 76 illustrations.

LONDON:
CHARLES GRIFFIN & COMPANY, LIMITED;
EXETER STREET, STRAND.

1912.

[All Rights Reserved.]

PREFACE.

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 Smelting—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 admirable 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 Journal, 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 Machinery 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 work.

University of Birmingham,
May, 1912.

CONTENTS.

LIST OF ILLUSTRATIONS.

TABLES.

COPPER SMELTING.

LECTURE I.

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 earliest 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 usefulness 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 characteristic in prehistoric times, that archæologists assign to one of the epochs the name of the Bronze Age. As is well known, archæological 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 development. 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 commencement 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 Nehosheth, from the root Nahásh, to glisten. This was translated as χαλκὸς (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 as 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 Cyprus, 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 ♀. 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 Boyle 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 Phœnicians 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 concluded 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 earliest establishments was situated at Neath—a fact recorded in a publication of 1602. The works probably existed for a century before that date, and the copper smelters at Swansea were established about 120 years afterwards.

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 wood 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 successive 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 establishing 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 provision 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 considerable 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 importations 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 abundant 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 was 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 with 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 Japan probably ranked next to Great Britain as producers, small amounts of ore from Austria, Scandinavia, and Italy made up the remainder. Thus at the commencement 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 development of Cornish copper mining that the use of steam power in engineering 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 successful 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 development 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, moreover, 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 simultaneous 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, the standard price of the metal being subsequently raised. Mine-owners commenced to seek for a remedy, their ultimate endeavour being to substitute, for the exportation of their ores, smelting operations at or near the mines themselves. In 1842 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 proportions of ore 21 per cent., regulus 38 per cent., and bars 40 per cent., they subsequently became ore 1½ per cent., regulus 3½ 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 obliged 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 Swansea. 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 supply diminishing with considerable rapidity.

In 1866 the great Calumet and Hecla mine at Lake Superior commenced 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 mining 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, many 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 probability 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. When, 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 processes.

As will be explained later, copper smelting of sulphide ores is essentially 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 successive roasting stages, for which purpose oxidised materials, such as roasted or oxidised ores, were added to the charge.

The Best-Selecting process, and the Nicholl 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 Brückner cylinder); not, however, adopted for copper smelting till many years afterwards. Later—Arrangements for using roaster gases for sulphuric acid manufacture.

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

(c) In Blast-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 smelting.)

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 difficulties, 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 reverberatory 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.
• (6) 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 the mines, and with increased consumption, stimulated by the growth of the electrical industry, the demand for metal increased so quickly that developments naturally 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, with 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 development 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 modern 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 considerably 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 1850–1860 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 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 combine.

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

In 1899 the Amalgamated Copper Company was formed in the United States. This corporation was established 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 expensive law suits, and the price of the metal dropped again with great rapidity. Trade subsequently revived and expanded, the consumption of copper increased and appeared to overtake the rate of production, whilst stocks diminished and the price advanced, until, 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 comparatively 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 13½ 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 marketing 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 with regard to the cost of production estimated at different plants. Of the American output of about 480,000 tons in 1909—

Copper Statistics.—The outstanding features which attract attention in the statistics of copper production will 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 I.:—

TABLE I.—THE PRODUCTION OF COPPER
(Short Tons of 2,000 lbs.).

Fig. 2.—Annual Production of Copper.

In Table II. is indicated the distribution of the American production among the various States.

TABLE II.—North American Production of Copper
(in Short Tons of 2,000 lbs.).

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 curtailment 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.

References.

Percy, John, “Metallurgy (Copper).”

Gowland, William, Presidential Address, Trans. Inst. Mining and Metallurgy, vol. xvi., 1906–7, pp. 265–291.

Stevens, H. J.,“The Copper Handbook.”

Brown, N., and Turnbull, C. C., “A Century of Copper.”

Engineering and Mining Journal, “Copper Production.” May 6th, 1910, p. 891.

Mineral Statistics of the United Kingdom.

Mineral Industry.

LECTURE II.

The Uses of Copper: as Metal and as Alloy—The Physical Properties of Copper—Effects of Impurities—Mechanical Properties—Chemical Properties.

The Uses of Copper.—Generally speaking, the industrial applications 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.
• (b) 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-conductivity copper.” The demand has, to a large extent, increased irrespective 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 engineer 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 very great purity. 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 wire, 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 transformed 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, 150 lbs. per mile.
• Minimum diameter, ·95½".
• Maximum diameter, ·98".[1]
• Minimum breaking strain, 490 lbs.
• Minimum number of twists, 25 in 3 inches.
• Wraps required, 6 times round wire of its own diameter, unwrapped, and again wrapped without breaking.
• Maximum resistance per mile at 60° F., 5·857 ω.

• B. Post Office Specification.
• Weight, 500 lbs. per mile.
• Minimum diameter, ·135¼".
• Maximum diameter, ·138¾".
• Minimum breaking strain, 950 lbs.
• Minimum number of twists, 30 in 6 inches.
• Wraps required, 6 times round wire of its own diameter, unwrapped, and again wrapped without breaking.
• Maximum resistance per mile at 60° F., 2·928 ω.

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

(b) Engineering 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 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 0·51 per cent.
• 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,43·36 per cent. of original length.
• Contraction of area, 45·9 per cent.
• Close bend test, Double.

The effect of temperature and 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 ductility. The diverse effects of different impurities on strength and ductility 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 like, 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.
• Coinage 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 manufacture.

The uses of the copper alloys may also be arranged in two classes—(a) engineering uses, and (b) general uses. Of the brasses, those containing 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 utilised 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, etc.

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 knowledge.

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 manufacturing 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.

Fig. 3.—Equilibrium Diagram, Cu-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 perhaps still greater importance is a knowledge of the influence exerted upon these properties by the circumstances which usually attend working practice; such as, for example, the various common impurities, 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.

TABLE III. —Influence of Impurities on the Electrical Conductivity.

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 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 purplish-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 appearance, 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 employment for heaters, steam-coils, and the like.

Fig. 4.—Influence of Arsenic and Antimony on the
Electrical Conductivity of Copper.

Power of Dissolving Gases.—When molten, especially under reducing conditions, the metal possesses the property, common to many others, of absorbing gases such as carbon monoxide, hydrogen, hydrocarbons, sulphur dioxide, etc., which are moreover, 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, appears to reduce the dissolving power of the liquid metal for these gases, or possibly to increase their solubility when the copper is solidifying, 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 characteristics of overpoled copper probably arise from this cause also.

Impurities[2] 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 Hampe’s 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 compounds 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 other workers with more delicate means of investigation at their disposal.

Oxygen in Copper.—Molten copper has the power of dissolving its oxide, Cu₂O. When the melted metal is exposed to oxygen, this oxide is produced and passes into solution in the liquid, yielding a series of binary alloys, of which the oxide acts as the second constituent. The equilibrium diagram of the series, as worked out by Heyn[3] ([see Fig. 5]), affords a good indication of these relationships, and throws light on several features connected with the presence of oxygen in copper.

Fig. 5.—Relations of Copper and Oxygen.

It will be observed that when molten oxygenated metal containing less than about 0·38 per cent. of oxygen solidifies, copper crystallises out first, whilst later, in between the copper crystals, there solidifies a eutectic of copper and cuprous oxide. This eutectic contains about 3·45 per cent. of cuprous oxide, equivalent to 0·38 per cent. of 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 Cu₂O : O :: 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 circumstances.

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.

ab

Fig. 6.—Microstructure of Copper containing Oxygen (Heyn).
a. Hypo-eutectic.b. Hyper-eutectic.

Oxygen 0·13 per cent. = 1·16 per cent. Cu₂O.
Oxygen 0·53 per cent. = 4·7 per cent. Cu₂O.

Fig. 7.—Relations of Copper and Arsenic.

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 compound produced is probably Cu₃As (28·3 per cent. of arsenic), which passes into solution in the excess of metal, and on solidification the copper retains this arsenide in solid solution. As in the case of all such solid solutions, the solidification takes place over a range of temperature represented between the liquidus and solidus curves; the purer metal crystallising out first, followed gradually by crystals of copper which become progressively richer and richer in arsenic (still in solid solution). In the case in question, diffusion of the arsenic throughout the crystalline mass proceeds but slowly, and as a result, the metal, as usually obtained in the cast state, shows fringes of such arsenic-rich copper. By annealing, diffusion is greatly assisted, and the material gradually becomes homogeneous, as is seen on microscopic examination. There appears further to be some decrease of this solubility with fall of temperature when the arsenic is high, leading sometimes to a separation of the arsenide itself at the crystal boundaries.

Antimony appears to form an analogous compound, Cu₃Sb, also capable of passing into solid solution in the copper, but to a rather smaller extent than the corresponding arsenide. The fringes are therefore more pronounced, and the decrease of the solubility on further cooling is also more marked.

Bismuth.—The influence of even minute quantities of bismuth on copper is notorious. Bismuth appears to be soluble in liquid copper, but not in the solid metal. In consequence, when copper containing bismuth solidifies, the copper crystals separate first, whilst the liquid bismuth still remains between them, until the metal reaches a temperature of about 268° C.—the melting point of bismuth—when it too solidifies in situ. The presence of such envelopes of very brittle, fusible, and limpid bismuth material explains much of the harmful effect of this impurity. These envelopes are found to consist almost entirely of practically pure bismuth. Oxygen converts the bismuth into a more compactly crystalline oxide, much less fusible and harmful. Arsenical copper tends to the scattering of the bismuth globules among the fringes which are formed during the gradual process of solidification over the range of temperature already indicated, and thus renders this impurity to some extent less dangerous.

Lead behaves in apparently much the same way as bismuth, and the effects produced upon it by the presence of oxygen and arsenic are probably similar.

Selenium and Tellurium probably exist in the form of selenides and tellurides, which are characterised by marked brittleness and fusibility.

Mechanical Properties of Copper.—The mechanical properties of commercial copper are influenced to a vital degree by the conditions associated with working practice, such as composition, previous mechanical and thermal treatment, temperature of working, etc. As has been already indicated, it is the possession by the copper of certain mechanical qualifications which leads to its employment by engineers, and it is, therefore, necessary to consider the influence of the above conditions, when reviewing the mechanical properties of the metal.

Much of the copper employed for general engineering work (apart from electrical and alloying purposes) is of the quality designated as “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 practically 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 considered, but attention must be directed to the actual condition and distribution of each constituent. Owing largely to the difficulties of determining 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 sufficiently concise to allow 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 circumstances of working practice may react through these to a considerable extent.

Many of the more general results have, however, long been known to engineers from practical working, 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 annealing. The annealing effect commences at about 300° C., but proceeds more effectively at higher temperatures, the factors of annealing temperature 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 se, 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 copper of 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 important influence on the results.

Hardness.—Pure copper is a comparatively soft metal. It is hardened by mechanical work—the hardness of rolled copper, determined 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 of more general standardisation than is at present possible. The tensile strength of pure cast copper is 8 to 9 tons per square inch. Mechanical 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 strengthening 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 in minute quantities, and especially on heating. Bismuth in the proportion of 0·005 per cent. lowers the malleability and ductility considerably, and recent reports state that 0·02 per cent. bismuth renders copper cold short, that 0·05 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 properties.

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, Cu₂O, is obtained, which is readily detached by quenching and hammering. Water at ordinary temperatures is without effect upon copper; concentrated sulphuric and nitric acids have little 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 provided 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.

References.

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

Proc. Inst. Mech. Eng., 1893; Dean, p. 139; Blount, p. 164; 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 Eng., 1908–9.

Hughes, G., “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. Soc., 1906.

Hiorns 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).

Zeitschrift für Berg. Hutten and Sal. Wesen,

Hampe 1873, xxi., 218; 1876, xxiv., 26

Chemiker Zeitung, 1892, No. 42, p. 16.

Reports, Royal Tech. Testing Institute.

Heyn, E., “Copper and Oxygen.” Charlottenburg, 1900, p. 315.

Metallographist, vol. vi., 1902, p. 48.

Arnold, Engineering, vol. lxi., p. 176. Feb. 7, 1896.

Roberts-Austen, Second Report, Alloys, Research Committee. Proc. Inst. Mech. Eng., April, 1893, p. 114.

Rudeloff, Mittheil. König. Tech. Versuchs. Anstalt., 1894, ii. (b), pp. 292–330; 1898. 16a, pp. 171–219.

Lawrie, Bull. Amer. Inst. Min. Eng., 1909, pp. 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.” Metallurgical and Chemical Engineering, 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). Eng. and Min. Journ., Jan. 20, 1912, p. 181.

LECTURE III.

Compounds of Copper—Copper Mattes—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 compounds are the oxides, the sulphides, and the silicates.

Copper Oxides.—Of the oxides, two are of importance—cuprous oxide, Cu₂O, 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 heating 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 heating with carbon, the metal being also obtained if the oxide be heated in the presence of reducing gases; it combines readily with silica when heated, yielding fusible silicates.

When cuprous oxide is heated with sulphide of iron, the copper, having a greater affinity 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 copper 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 Nicholl 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 Cu₂S and CuS, the former only 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—

	Cu2S + 2O ➡ 2Cu + SO2
	2Cu + O ➡ Cu2O

which take place simultaneously, the copper represented in the first equation being oxidised spontaneously according to the second, and the resultant is the reaction Cu₂S + 3O ➡ Cu₂O + SO₂.

In the furnace operations, some of the SO₂ 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 SO₂ 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 sulphide which is thus produced, has 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. They may contain from a mere trace to upwards of 80 per cent. of copper, and in ordinary work, sulphide of iron is the other constituent present in the greatest proportion, but sulphides 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 work. 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 widely differing results. Röntgen made an exhaustive investigation of the system FeS—Cu₂S, and published a very complete diagram of the series, working with FeS and Cu₂S 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 formulæ Cu₂S 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. 8. 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 80 per cent. matte.

The problem of the constitution of mattes is, however, a very complex one, and is not yet satisfactory settled. An interesting view was put forward by Gibb and Philp. Mattes corresponding to the formula 5Cu₂S . FeS (copper 71·7 per cent.), when examined microscopically, appeared to be homogeneous, and indicated some form of combination 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 quantity 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 solubility was lessened, and moss copper was set free in the solid matte.

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

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 the separation of moss copper is the partial decomposition of Cu₂S, 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 equilibrium.

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

Copper Silicate is formed by the action of copper oxide and silica on heating. The silicate is decomposed when heated in the presence of sulphides, resulting in the formation of sulphide of copper and silicate of the second metal, in consequence of the great affinity of copper and sulphur. Upon this fact depends the extraction of copper from various silicate ores, as well as the cleaning of slags high in copper, which are often added to the sulphide charges in the furnace with this object. When heated with iron, the silicate is reduced to metallic copper with the production of silicate of iron; it is also reduced by carbon in the presence of metallic oxides capable of uniting 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 which they are produced, or by special trade names. The most important variety is:—

Electrolytically-refined High-conductivity Copper, which is largely used for electrical work. The methods by which 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 gold. As ordinarily produced at the electrolytic refinery, it is in the form of cathode plates, often about 3 feet × 2 feet 6 inches by ¾ 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 are particularly pure, and on smelting and furnace-refining yield a metallic product of great purity which also possesses good mechanical properties. It is, therefore, particularly suitable for electrical work. By reason of its satisfactory properties, Lake copper realises prices which usually rule somewhat higher than those of ordinary electrolytic copper as quoted on the New York market.

Best Select Copper.—For the production of copper alloys, such 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 was employed for obtaining 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, rather than for obtaining specially pure copper. The product is essentially a British one, and was largely used for the manufacture of high quality alloys.

“Tough Pitch Copper.”—The operation of “bringing copper up to pitch” has for its object the imparting to the metal of the toughness and mechanical strength required for industrial service. The process resolves itself into the adjustment of the correct proportion of oxygen, the function of which is largely to eliminate the gases from the copper, or to overcome their deleterious effects, as well 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 refining 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 removable 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 imparting 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 per 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 oxidisable 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 oxidule-eutectic are seen on examination under the microscope. This material is known as Dry Copper; 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.

If the poling—that is, the reduction of the oxidised constituents of the tough pitch copper—be carried too far, the metal becomes brittle again, being 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 they arise from the removal of oxygen from the oxygenated constituents, and the withdrawal from the metal of the protecting influence 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 compounds, 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 commercially useless as such, and is really an intermediate product, the metal being brought to pitch again by further aëration 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 commercial varieties of copper are:—

• Electrolytically-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 put on the market direct in this form. Being produced from fairly pure ores, which carry but little silver and gold, the converter metal may be sufficiently pure to render electrolytic refining unnecessary, 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.

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

Cement Copper is produced by wet processes, usually by precipitation from copper-bearing solutions by means of iron, the product being a rather impure reddish-brown spongy mass. Many varieties contain arsenic. It requires melting and subsequent refining to adapt it for service.

Blister Copper was the name given to the crude metal from the older type of furnace operations. Such copper contained large quantities of gas, particularly SO₂, which, tending to escape at the moment of solidification in the mould, gave a blistered 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 copper exhibiting similar features.

Chili Bar is an impure copper imported from Chili for refining. The composition varies, the metal usually containing 96 to 98 per cent. of copper, with indefinite quantities, sometimes small, of undesirable impurities.

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 accompanied by very varied gangue, according to the locality in which they are found.

They are best classified under three groups:—

• (1) Native Ores.
• (2) Sulphide Ores.
• (3) Oxide Ores.

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

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

TABLE IV.—Analyses of Various Commercial Coppers.

(2) Sulphide ores supply the bulk of the world’s copper, constituting upwards of 80 per cent. of the total.

(3) The oxidised ores are found in most copper districts, though 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 rapidly, 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 exceptional 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.

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

(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 Cu₂S. Fe₂S₃ (Cu 34·4, Fe 30·5, and S 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 pyrrhotite. It occurs principally in the older crystalline rocks, 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 5 to 5½ per cent. (with gold and silver valued at about £11 per ton of copper), whilst the Arizona ores vary, being often rich.

Chalcocite (also known as copper glance or redruthite) is much less important. The copper contents are 79·8 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:—

Bornite (Erubescite, Peacock copper ore), 3Cu₂S. Fe₂S₃, occurring in Cornwall, which passes with depth into chalcopyrite.

Tetrahedrite (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)₂, 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 sulphide ores, it is an extremely useful material to mix in the charge, as it supplies oxygen as well as copper. Malachite is still an important source of the metal in Mexico, Chili, and Bolivia, though not quite so much so as formerly, whilst it is specially important in the Tanganyika (Katanga) deposits, of which it constitutes the greater portion so far developed.

Cuprite, Cu₂O, contains 88·8 per cent. copper, when pure. It is widely 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 was 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 below.

Other oxidised ores include—

Azurite, 2CuCO₃. Cu(OH)₂, and Atacamite, CuCl₃. 3Cu(OH)₂, from Chili.

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 preparatory to smelting includes the processes of sampling, wet concentration, 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 with 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 departments 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 sampling 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 of the best practice, that it may be reviewed in brief, as an example.

Fig. 9.—Outline of Sampling Scheme, Anaconda.

Fig. 10.—Section through Sampling Mill.

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. Each section consists of a set of four sampling machines with intermediate 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 (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 one-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-chute.

Fig. 11.—Brunton Sampler.

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 through a 1-foot riddle of 100 mesh wire cloth, the very small quantity of coarser stuff remaining, is bucked down and added, and the whole is then thoroughly mixed in a canister of 1 foot side gripped at opposite corners, and rotated mechanically.

References.

Constitution of Copper Mattes.

Keller. Mineral Industry, vol. ix., 1900, p. 240. “Elimination of Impurities from Copper Mattes.”

Röntgen. Metallurgie, vol. iii., 1906, p. 479.

Hofman, Caypless, and Harrington. Trans. Amer. Inst. Min. Eng., vol. xxxviii., 1908, pp. 142–153.

Gibb and Philp. Trans. Amer. Inst. Min. Eng., vol. xxxvi., 1906, p. 665.

Heyn and Bauer. Metallurgie, vol. iii., 1906, p. 84.

Fulton and Goodner. Trans. Amer. Inst. Min. Eng., vol. xxxix., 1908, pp. 584–620.

Refining of Copper.

H. O. Hofman, R. Hayden, and H. B. Hallowell, “A Study in the Refining and Overpoling 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, Analyse and Eigenschaften des Kupfers.” Berg. and Hüttenmännische Zeitung, 1889, vol. xlviii., pp. 323–4; 1890, vol. xlix., p. 399; 1893, vol. lii., p. 19; 1901, vol. lx., pp. 77–79.

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

Sampling.

D. W. Brunton, “Modern Practice in Ore Sampling.” Mining and Scient. Press, Oct. 30, 1909. “Theory and Practice of Ore Sampling.” Trans. Amer. Inst. Min. Eng., vol. xxv., p. 826.

LECTURE IV.

Modern Copper Smelting Practice—Preliminary Treatment 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 matte containing from 40 to 50 per cent. of copper, which is then bessemerised.

There are however proceeding at present (owing to the successful working of basic-lined converters) developments which indicate that such 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 equally 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-milling.

Oxide Ores are rarely wet-dressed. They present much difficulty in treatment on account of their comparatively low density, which makes 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, (b) the concentration of the copper desired in the first smelting operation, and (c) the smelting method and furnaces adopted.

Wet concentration is only profitable when 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 circumstances 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 5½ 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 expensive 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 preliminary manipulation of the fine concentrates that are produced is rendered necessary. Such modified dressing schemes involve a maximum of coarse breaking and screening, the crushing and separating stages 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 1¼-inch and ⅞-inch feed.
• Middlings crushing in rolls.
• Middlings sizing by trommels.
• Middlings separation on fine jigs of 7, 5, 2½, 1½, 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 in 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, 1¼, ⅞ (and ⅜) inch size, are smelted in the blast furnaces.
• Fine Concentrates, 7, 5, 2½, 1½, 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.

Fig. 12.—Outline of Smelting Scheme at the Anaconda Smelter, Montana, U.S.A.

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 quantity 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.[4]

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 preparation.

Blast furnaces have many advantages which lead to their extended use in copper smelting practice, but one important feature, which also applies 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 be 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 ¼ to ⅜ 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,
• (b) Irregular working of the furnace and the charge,
• (c) Clotting and low 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 preliminary in any scheme of treatment involving the employment of the blast furnace on such material. Agglomerating is usually performed by one of two methods—(1) briquetting, (2) sintering. Of these, briquetting has hitherto been in very general use, but several advantages connected with the sintering process and the resulting product are leading to its adoption with much success in several localities, and attracting for it considerable attention at present.

(a) Briquetting.—Among the advantages of briquetting is the fact that it utilises large quantities of the copper-bearing slime produced at the concentrating plant, this material often possessing good binding properties which render it very suitable for briquette-making.

Fig. 13.—Sketch Plan of Briquetting Plant.

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

Fig. 15.—Chambers’ Briquette-making Machine.

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 (< ⅜-inch size), fine concentrate from the dressing plant, and coke (which is recovered from the reverberatory furnace gratings). The quantities used daily are somewhat as follows, though they are naturally subject to some variation, depending on supplies:—

and the composition of the briquettes is about—

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 up 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 × 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 employed.

(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 siliceous.

Charge.—Must contain 15 to 35 per cent. silica, 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 ⅝-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 1s. more per ton than for ordinary roasting.

Though this particular process is only, to the author’s knowledge, employed at a few 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 particularly suited for the class of material indicated above. The advantages 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 reverberatory furnaces.

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

(c) 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 realised 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 ¼ inch, and not more than 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.

Fig. 16.—Dwight-Lloyd Sintering Machine.

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 very important preliminary stage in the scheme of treatment of copper ores. It was formerly considered 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 operation.

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 subject.

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—

The copper takes up sufficient sulphur to form Cu₂S; the remaining sulphur combines with any iron which is available, forming FeS. These two sulphides, dissolving in all proportions, constitute the matte product of smelting.

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.[5]

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 concentration 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—Roasting.

(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.
• (b) 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.

(4) 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 1s. 8d. per ton of ore (equivalent to ½d. 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 liberty £34,000, which had been tied up in this manner.

Advantages of Roasting.—Illustrative of the conditions under which 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 sulphides, 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 considerable quantity of fine material is unavoidably produced, for which the most convenient treatment in such large quantities, under prevailing conditions, is in the reverberatory furnace. The atmosphere of this type 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.[6] 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 remelting operation, and so affords exact control of the concentration effected.

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

In modern practice the work of the reverberatory plant is controlled 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 exposing of a substance to the effects of heat and air, in order to oxidise it, and to render it more suitable for subsequent smelting operations.[7]

In the case of the ordinary sulphide copper ores, roasting not only (a) reduces sulphur, and so ensures good concentration on smelting, but (b) by oxidising the iron, provides a ready flux for siliceous gangues. 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—

FeS₂ ➡ FeS + S.

Iron Sulphide.—Sulphur has a great affinity for oxygen, to form SO₂ and it may be assumed that this reaction first takes place thus—

FeS + O₂ ➡ (Fe) + SO₂ (i.)

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

(Fe) + O ➡ FeO (?)   (ii.)

Or, combining (i.) and (ii.)—

FeS + 3,O ➡ 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 itself.

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

	2FeO + O ➡ Fe2O3.
	3FeO + O ➡ Fe3O4.

The SO₂ in the presence of oxygen and in contact with strongly 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 oxidation of the nascent copper—

	Cu2S + O2 ➡ (2Cu) + SO2
	(2Cu) + O ➡ Cu2O,

thus

Cu₂S + 3.O ➡ Cu₂O + SO₂;

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

In addition to the tendency to melt, copper sulphide roasts less perfectly 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 Cu₂S. Fe₂S₃, and accompanied usually by a large excess of FeS₂, it behaves very much like a mixture of these sulphides when treated in the roaster furnace, hence the reactions on roasting follow on the lines just indicated.

In practice the roasting is never carried to such a degree that all the sulphur is eliminated, since it is essential to retain some sulphur in order to collect the copper in the form of matte, and also because the time, and the cost of the fuel required to roast all of it off, would be prohibitive. Consequently, the products from the roasting of chalcopyrite consist principally of oxides of iron and copper, together with a certain amount of copper sulphate, very little iron sulphate, and some undecomposed sulphides.

The actual form in which the sulphur is present at the end of the roasting operation is not usually of very special importance in practice, especially where the previous experience with the roasted material determines the extent to which the roasting is conducted, since the greater part of the sulphur eventually produces the sulphide and constitutes the matte, on smelting the roasted charge; although some is also eliminated as SO₂ by interaction with oxides. In modern roasting practice, therefore, all that is usually required is to roast the ore down to, say, 5 per cent., 6 per cent., 8 per cent., or whatever proportion of sulphur is necessary to yield the required grade of converter-matte in the reverberatories, as judged by previous experience of the furnace plant and working. Much SO₂ 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 + ZnSO₄ + SO₂

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

Lead Sulphide 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 oxidation are largely analogous to those for other sulphides.

	PbS + O2 ➡ Pb + SO2
	Pb + O ➡ PbO

or,

PbS + 3.O ➡ PbO + SO₂.

Also,

2PbO + SO₃ ➡ PbSO₄.PbO (basic sulphate).

Arsenides are partly left as the corresponding oxides, whilst some As₄O₆ is evolved, and some basic arsenate generally remains.

Roasting Practice.

Favourable Conditions for Successful Roasting.

(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 continually 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 condition 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 class 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.—In modern copper-smelting work, the practice of roasting lump ores is practically obsolete. The conditions under which its use might still be justified are those associated with newer mining districts, where rapid concentration of heavy sulphide ore into matte is required, before the time is ripe for smelting the material pyritically, and where further, it is desired to employ the blast furnace for the smelting operations under these circumstances.

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 immediately suitable for blast-furnace work.

(3) 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 conditions are exceptional.

Amongst the many grave objections to open-air roasting are—

(a) It is very slow, since a long period of time is required for the oxidising effect to penetrate through massive lumps of ore.

(b) A large amount of capital is tied up in the material at the roast-yards.

(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-heaps.

(e) Difficulties arise owing to damage by the fume, and from interference by litigation.

There is one special instance of a modern smelter making a great success of heap-roasting—namely, at Rio Tinto—but the circumstances are peculiar, as the roasting is followed by leaching operations of the immense ore heaps in situ.

This branch of roasting need not be considered at length, and the older standard text-books give full descriptions of the various methods employed. The following particulars are important, however, when under exceptional circumstances such work has to be undertaken:—

The maximum and best average size under ordinary conditions is 40 feet by 24 feet, by 7 feet high above the bed of fuel. The height is important, and varies with the quantity of sulphur in the ore. The lower the sulphur content, the higher the pile; with about 40 per cent. sulphur, the best height is 6 to 7 feet; with 15 per cent. of sulphur, up to 9 feet; and if still less sulphur be present, the height may even be a little greater. Such a heap holds about 240 tons, and if the 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.

(b) 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 conveniently be precipitated on scrap iron. Enormous losses of copper may occur if these precautions are not observed; thus, at one period in the old roasting process in Tennessee, as much as 34 per cent. of the copper in the heaps was lost in 186 days.

Preparing of the Floor.—Remove roots and subsoil, fill space with 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 does 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 concentrates) 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 ⅜-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 required.

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.
• (b) By furrowing and rabbling the charge.
• (2) Continually to expose fresh surfaces of ore to oxidation.
• (a) First by hand-rabbling.
• (b) 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 × 13 feet 6 inches, constructed of firebricks set on end, and with a fire-box 7 feet × 2 feet 3 inches × 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 labour. 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 about 2 inches at a time, thus ensuring better control of working. By this means the best type of hand-calciner was arrived at, consisting of four beds, each 16 feet × 16 feet, the whole charge being moved forward from one bed to the next at each stage of the process.

In 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.
• (b) 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 discharge, 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 × 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 breakdowns, and the subsequent modifications of this form consisted largely of devices for the purpose of overcoming such difficulties.

Modifications 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.

Brown (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 improvements 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 runs 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 is about 105 feet long × 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 above, except that the hearth is bent round in order to save space.

Pearse-Turrett (1892 at Argo).—In this type of furnace the bed is curved round in the form of a circle. The rabbling ploughs are carried 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 employed as fuel.

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

(a) Intermittent Working—The Brückner 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 introduced 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 Brückner 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 considerable 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—(b) 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 materials.

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

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

White (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 work 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, inclined 6 inches in 60 feet, makes 8 revolutions per hour, and has a capacity of 10 tons of charge per day.

Argall 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 roaster furnace, and that most generally employed, is the MacDougal type. The first furnace on this principle was invented by Parkes in 1860. The design embodied two hearths, one above the other. Vertically 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 works.

Principles of the MacDougal Type.—The furnace consists of an iron cylinder lined with brick. Six circular hearths are constructed inside, one above the other, and the vertical spindle carrying the arms and ploughs for each hearth passes through the centre of the furnace. The ore is ploughed towards openings on each hearth, which communicate with the hearth below; the charge thus travelling from the outer 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 12 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 easily 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.

Fig. 19.—MacDougal Roaster—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, where the rapid wearing out of working parts, and the difficulty of their removal, repair, and renewal interfered greatly with efficient working.

Many of these troubles have now been overcome in the Evans-Klepetko type, and in the still further 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 centres.

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 ploughs simplifies the mechanism.

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

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

(5) Very little fuel is required, none with heavy sulphides (except for 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 by adjusting the rate of passage of the ore through the furnace, which is regulated by varying the ore feed and the speed of rotation of the rabbles.

(8) Great saving in labour costs and difficulties. The labour in roasting plants is extremely arduous, on account of the high temperature of the material, and is dangerous on account of the atmosphere.

The Evans-Klepetko-MacDougal Roasting Furnace Plant at Anaconda.—The roasting plant at Anaconda formerly consisted of 56 Brückner 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 Brückners by MacDougal roasters is reckoned at about 5 cents (2½d.) 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.

Fig. 21.—Spindle Connections
and Guide Shields of
Evans-Klepetko Roasters.

Details of Furnace.—Height, 18 feet 3½ 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 1½ 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 resulting from one set of blades are turned over by the other. The blades are set so as to direct the ore from the outer to the inner edge or vice versâ, according to the particular hearth. The spindle and connections are protected from falling ore by shields which are bolted on. The rabbles move slowly, making a 2½-inch furrow in a 5-inch layer of material.

Capacity.—40 to 45 tons per day each, reducing the sulphur in the charge from 30 per cent. to about 8·0 per cent. The output of the plant is about 3,000 tons of calcines daily.

System of Working.—Since reverberatory furnaces are used essentially 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 settling tanks, and screened lime-rock flux (too fine to be used in the blast furnaces). The limestone lightens 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, 1¼ tons of lime-rock, and 1¼ 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.

Fig. 22.—Rabble-blades and Bases.

1st Hearth.—Temperature about 230° C. (black heat). This is 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.

3rd 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.

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

5th 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 SO₂ by volume, 5 per cent. by weight. The ore takes 2¼ 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 foreman 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 which 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 fireplace. Here the output per furnace per day approaches 55 tons, roasting the sulphur from 20 per cent. down to 10 to 11 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 prevailing at this particular camp.

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

whilst the product (“calcines”) has an average composition of

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

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:—

References.

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. O., “Recent Progress in Blast Roasting.” Bulletin Amer. Inst. Min. Eng., No. 42, June, 1910.

Austin, L. S., “The Washoe Plant of the Anaconda Copper Mining Company.” Trans. Amer. 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.

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.

LECTURE V.
Reverberatory Smelting Practice.

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

The Functions of the Reverberatory Furnace. —The reverberatory 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 separate thoroughly.

In spite of the neutral atmosphere, however, the smelting of the roasted materials usually results in a higher concentration than would be expected from the calculation of the sulphur, copper, and iron in the charge. The reason of this is that the smelting operation results in some further elimination of the sulphur, which causes the production of a higher grade matte. This additional elimination of sulphur in 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 + SO₂
Cu₂S + CuSO₄ ➡ 3Cu + 2SO₂,

which indicate a further addition of copper to the matte, and a corresponding loss of sulphur. Thus a typical reverberatory charge of the following composition:—

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

[8]Cu (8·3) ➡ Cu2S  10·4				Cu 8·3				Cu 30	per cent.
		=		S 8·4		or		S 30	"
S (8·4−2·1) ➡ FeS  17·6				Fe 11·3				Fe 40	"

In actual practice however, the matte resulting from the reverberatory smelting of the charge had the composition—

	Cu	45	per cent.
	S	27	"
	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 a knowledge of the composition of the roaster product, to regulate and control the grade of the matte produced at the reverberatories. In 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 simply charges what is sent him from the roasters, and practically nothing else is put in,[9] 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 satisfactory, 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 day.

Reverberatory smelting is essentially a British process, developed in Wales, as already explained, owing to a plentiful supply of good furnace coal yielding a long flame, and also of good refractory material. Many Swansea workmen were, in the early days of American development, and are still, employed in charge of such copper furnaces, and it is largely 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 reverberatory 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 conducted.

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 the methods of achieving the proper conditions can best be stated as 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 quantity 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 methods 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 avoid 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 withdrawal 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, and improvements in the economy of the process were in the direction of 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 utilise 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 rapidity 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 charging.

The advantages of large grate area therefore include:—

• (a) Much less cooling of the furnace by frequent additions of fuel.
• (b) 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 is melted chiefly by the heat from the hot gases passing over it, and 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 fuel in the fire-box, and by the drawing of the 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 drawn 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 by 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—

• (a) Special power and machinery are required.
• (b) The intense action near the fire-bars produces, from the ash of the coal, a massive clinker in a semi-fused condition, difficult to deal with.
• (c) It is stated by smelters to have a cooling action near the fire-bridge.

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

The addition of fresh coal to the fire causes the production of large quantities of volatile hydrocarbons which require an increased air supply for proper combustion, and this air admission is just prevented by the blanketing action of the fresh fuel added. This is indicated by the red smoky flame, and means waste and cooling. The difficulty is overcome by the arranging of a series of air-holes 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 1½ 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 off-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 half-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.
• (b) 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 dealing 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.

(b) 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 decreased.
• (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. Rapidity of melting is an indispensable feature of modern work. The conditions necessary for rapid melting have been reviewed above.

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 metallurgist) 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.—Development in Size of the Reverberatory Furnace.

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 fuel is burned, and the enlarging of the fire-box had a specially important influence in effecting this rapidity of combustion.

Then, with the size of grate fixed and the most efficient burning of 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.
• (b) 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 by increasing the length, and this is limited only by the distance from the fire-box to which the flame can maintain the temperature necessary for keeping the charge in a state of perfect fluidity. For many years the length was regarded as limited to 50 feet, smelting about 2·7 to 3·0 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, 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.

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

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) Settling 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 wherever 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 reverberatory 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 modern 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 the 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 temperature.

(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 facilitated. 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 maybe 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 half-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 device.

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 with 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:—

The draft record is placed close to the charging 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 charging of hot “calcines,” and the continuous or regular withdrawal of slag and of matte when required.

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

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 siliceous bottom are slagged off before they get an opportunity of reaching it. The hearth bottom, if properly put in, is practically permanent.

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 silica bricks. The employment of higher temperatures in modern 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 hoppers situated in the roof, near to the walls, has lately been introduced with a view to protecting the furnace sides from the corrosive action of the slag, and to exposing a suitable siliceous flux to this material. This appears to have fulfilled its purpose to some extent, but various difficulties have been encountered in practice, especially 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 trouble.

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 furnace.

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 siliceous 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. 25.—Skimming Reverberatory Furnace, Anaconda.

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 as 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.

Fig. 27.—Reverberatory Furnace under Construction.

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 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 silica bricks, and upon this, the actual working 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 layers.

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 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 is unaffected 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 little wear, the beds became raised in layers, this being 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 liquid 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 weaknesses, 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 by putting in the whole layer of 26 inches of sand at once, and firing as hard 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 off the bed after two years’ working.

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

Large Reverberatory Furnaces: Details of Construction.—The large furnaces at Anaconda were the first of the modern type to be constructed, they have met with enormous success in practice and constitute the standard form. Similar furnaces are now in operation or under construction at many of the large modern camps, and are of similar design and construction.

The hearth 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 ¼-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 through.

Working of the Reverberatory 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 carries 5 tons of charge, travels from the roasters and enters the reverberatory building by an overhead track running above the charge bins of the furnaces. It discharges through hoppers into the bins which extend across the entire width of the hearth. Bins were formerly arranged at intervals all the way down the furnace, but now only the two bins nearest to the fire-bridge are employed. Into the back bin, 10 tons of charge are placed, and into the other, 5 tons. Each of these bins discharges through two hopper discharge openings, feeding the furnace through holes in the roof (Figs. 29, 30), which are closed, when not in use, by round firebrick tiles 20 inches in diameter and 2½ inches thick; these are moved in and out of position by means of levers operated from the fire-box platform.

The temperature maintained in the furnace is high, approximating to 1,500° C., and just previous to dropping in a fresh charge, a workman, by means of a rabble, feels about the hearth below the charging hopper in order to ensure that all of the previous charge has been melted, and that none of it is sticking to the furnace hearth. By employing only the comparatively small quantities of 15 tons, this sticking is avoided, since such charges are not heavy enough to sink unmelted through the 8 inches of slag and 8 inches of matte in the furnace. The former practice of feeding charges amounting to 45 tons through hoppers situated all the way along the furnace had given serious trouble in that respect, and had consequently to be discarded. When the examination of the hearth is completed, the time occupied being very short, the side door is closed, and sealed with sand; the covers to the holes in the roof are now withdrawn, the gates closing the hoppers pulled back, and first the 5-ton, then the 10-ton charge is dropped into the furnace. The whole operation, including the preliminary opening of the door to test the furnace bottom, occupies five minutes.

Very little 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 platform, for the skimming 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.

As soon as the charge has been dropped on to the pool of molten material, the mass appears to spread out over the surface and float towards the skimming door, in a thin slow-moving stream which disappears when about half-way down, being usually melted within one hour. The former 40-ton charges required as much as eight hours for melting.

Owing to the great heating effect of the large bath of hot material below, and of the intense flame above, there is but little cooling action on adding the fresh charge; whilst with this length of furnace, practically all the dust is settled, and very little is carried into the flues.

Coaling.—The quantity of coal employed amounts to 20 to 25 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.

Fig. 29.—Fire-box End of Reverberatory Furnace, showing massive
Bracing, Charge Bins, and Charging Levers—Anaconda.

Fig. 30.—Interior of Reverberatory Furnace (looking towards Skimming Door),
showing Expansion Spaces in Roof, and Charging Holes—Anaconda.

Coal is charged every 40 minutes in quantities of 1½ tons at a time, from bins which extend across the entire width of the fireplace, 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 Diamondsville, 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 4½ to 6-inch centres, and is maintained at a depth of about 27 inches. Grating 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 quantities are dropped. The material all falls down a bank inclined at 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 the fuel charged into the furnace is recovered in a useful form. This coke is used up as a constituent of the briquettes.

TABLE VI.—Daily Report—Reverberatory Furnaces.
August 17th, 1908 (Good Day).

			Delays.
Furnace No.	Copper Material Smelted per Ton of Coal	Cost of Coal per Ton of Metal Melted	Waiting for Coal	Waiting for Calcines	Miscellaneous	Total Delays	Boilers Working	Ladles of Matte in Furnace at End of Day.
	Tons	$	Hours	Hours	Hours	Hours	Hours
1	4·77	0·95		—— No delays. ——			24	10
2	4·85	0·94					24	10
3	4·47	1·02					24	10
4	4·61	0·99					24	10
5	4·29	1·06					24	10
6	4·77	0·95					24	10
7	..	..					..	..
8	4·85	0·94					24	10
Total	..	..	..	..	..	..	168	70

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’ demands, the matte is tapped and run outside the reverberatory building, being cast into large matte-beds. The tap-holes are situated between the second and third doors, and between the fourth and fifth; and each consists essentially of a copper plate 2 inches thick and 25 inches square, which at first stands back 9 inches from the outside of the wall. Through this plate a 1-inch hole has been 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 months.

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.

Composition of calcines		SiO2,	29·5	per cent.
		FeO,	37·3	"
		S,	7·7	"
		CaO,	2·7	"
		Copper,	8·6	"
Composition of slag,		SiO2,	39·4	per cent.
		FeO,	40·7	"
		CaO,	4·3	"
Copper in matte,			38·6	"

TABLE VIII.—Monthly Report—Reverberatory Furnaces.
Total Charge—All Furnaces.

			Analysis.
Slag Calculation─			Calculated.	Actual.
SiO2	in slag,	14,569 ÷ 38,538	37·8	37·1
FeO	"	16,214 ÷ 38,538	42·1	43·2
CaO	"	1,242 ÷ 38,538	3·2	2·8
		———————	———	———
		32,025 at 83·17 = 38,538	83·10	83·10
		══════════	════	════

Fuels for Reverberatory Furnace Work.—The chief requirements of the fuel for good reverberatory work will now be apparent, particularly with regard to length of flame. This depends to a large 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 working.

The device of using pulverised coal as a fuel has attracted attention 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 preventing 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 blanket which retarded the melting of the material by the flames. The method does not appear at present to offer much promise of extended application to copper smelting.

Oil Fuel in Reverberatory Practice.—The successful application of oil as a fuel marks a useful advance in reverberatory practice, particularly in connection with the working of large furnaces.

On several of the smaller plants, oil fuel has been in use with considerable 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 preliminary difficulties appear to have been to a large extent successfully overcome. The work at the Cananea Smelter with oil fuel, and the discussion on Ricketts’ first report of his experience, afford valuable indications of the possibilities of this method. Working on charges consisting to a large extent of flue-dust, several thousand tons of 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 particularly 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.

Fig. 31.—Shelby Oil-burner for Reverberatory 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 temperature 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 burning 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 lbs.) 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 described.

Costs of Oil-fired Reverberatory Working.—Ricketts has contributed 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 to July, 1911, inclusive.

Furnace Days, 312·5.

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 occurrence 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 particular 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 reverberatory 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 charge.

It is usual to smelt in the reverberatory furnaces, where such are available, the greater portion of the dust which accumulates in very 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 temperature 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 prevailing temperature would otherwise be sufficient to cause fusion. The particles of flue-dust moreover, 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 above the charge and in the flues. Such oxides clot only with the greatest difficulty, and are characterised by comparative infusibility and poor conducting power, and hence are found to melt with considerable difficulty when treated in the reverberatory furnace.[10]

Roasted fine concentrate, on the other hand, constitutes an ideal material for the reverberatory furnace charge, and the system of passing both the concentrate and the flux through the roasters has 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 would otherwise interfere with 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 when 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 reverberatory 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 proportion of the reverberatory charge, working in with 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 treatment, but also on account of the actual losses in the economy of the furnace processes and the cost of rehandling, etc. In modern 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 will be largely overcome. The smelting of fine concentrate in the blast furnace has up to the present been considered judicious where circumstances have rendered imperative 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 developments 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 values in the sintered product, as suggested by Peters, the difficulties in connection with excessive 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 reverberatories.

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 material.

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 blast-roasting.

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 of fine concentrate higher in silica and lower in iron and sulphur contents, which is not quite so suitable for ordinary roasting (owing to the necessity for external heating, due to lower fuel values) is eminently suited for blast roasting or sintering processes, yielding lump products very suitable for subsequent blast-furnace treatment.

The reverberatory furnace thus deals most successfully with fine table concentrates high in iron and sulphur, moderately low in silica; roasted, with its required flux, to the necessary extent, and then charged whilst still red hot into the furnaces. To relieve the reverberatories 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 treatment.

References.

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. 1234–1236.

Ricketts, L. D., “Experiments in Reverberatory Practice at Cananea, Mexico,” and discussion, Trans. Inst. Min. and Met., vol. xix., 1909–10, pp. 147–185.

Ricketts, L. D., “Developments of Cananea Practice.” Engineering and Mining Journal, Oct. 7th, 1911, p. 693.

LECTURE VI.
Blast-Furnace Practice.

Functions of the Furnace—As Melting Agent—Reduction Smelting—Oxidation in the Furnace—The Pyritic Principle—Features of Modern Practice: Water-Jacketing, Increase in Furnace Size, External Settling—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 range 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 smelted in the charge, the directly reducing function was utilised to a very much smaller extent. In the reducing atmosphere then maintained inside the furnace, the sulphides liquated and melted down 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 interactions 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 introduced with a view to increasing the concentration of the copper, this being 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. Reverberatory 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 preheated 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 construction, 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.

(f) 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 character.

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 effected.

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 quantity by themselves to be worked by this method, which involves also 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, however, of interest in so far as they apply to certain analogous problems in modern work. These difficulties in reduction smelting arose largely from three causes:—

• (a) Losses of copper in the slag.
• (b) Simultaneous reduction of iron with the copper.
• (c) Chilling 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 either

• (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 converter 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 application, 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 settling, promote copper losses from the other causes noted above.

Moreover, the high melting point of the metal and its great conductivity 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 difficult 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 constituents 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 constituents 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 itself.

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 principle of smelting).

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 + Cu₂S ➡ 4Cu + SO₂

2Cu₂O + Cu₂S ➡ 6Cu + SO₂

CuSO₄ + Cu₂S ➡ 3Cu + 2SO₂

whereby copper is produced and sulphur is eliminated as SO₂. 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 + xFeS ➡ Cu₂S . xFeS (matte) + Fe (oxidised and enters slag).

Copper silicates readily interact with iron sulphides in the charge, producing copper sulphides and iron silicates, thus—

Cu₂O . xSiO₂ + FeS ➡ Cu₂S (enters matte) + FeO . xSiO₂ (enters slag).

6(CuO . xSiO₂) + 4FeS ➡ 3Cu₂S (enters matte) + 4(FeO . xSiO₂) + 2xSiO₂ (enters slag). + SO₂.

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 favourable 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 laboratory scale, the other from actual industrial practice.

Starting from theoretical considerations, John Holway demonstrated 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 charge are strongly marked ([see Table IX., p. 120]).

TABLE IX.—Effect on Coke Consumption of
Increased Sulphur in the Furnace Charge
(Keswick Smelter, Cal.).

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 occasioning 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 ⅜ 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 11 to 12 per cent., the coke consumption has been steadily reduced until it now amounts to about 9 per cent. only.

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 supplementing the heat derived from the sulphide fuel, and the proportion employed in modern work is reduced to the lowest possible quantity. Not 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, since, 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 would 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 converter 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 proportions. 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.

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

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 gradually 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, 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 1½ 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 excessive 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 applied to certain branches of cast-iron refining, and in 1875 the Piltz 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 of water-jacketing was accomplished. This feature in blast-furnace work was rapidly and very successfully developed by the American copper smelters when the new establishments in the West were opened up, and the substitution of the older form of lining by metallic water-cooled jackets, which in comparison are practically indestructible, immediately led to an enormous improvement in smelting practice.

The modern blast furnace is essentially a water-jacketed shell from charging floor to base plate, rectangular in plan, and completely sectionised.

Many of the advantages of such a furnace construction are apparent, and have been referred to in discussing the furnace as a melting agent. The salient features of the modern water-jacketed furnace 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 water-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 marked.

(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 scarcity 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 incites internally at the tuyere level, the capacity being about 4 tons per day. Furnaces at this period were usually square or circular in section.

Fig. 33.—Blast Furnaces under Construction, showing Fixing of Jackets, Bottom Plate,
Method of Support, Sectioning, etc. (T. E. Co.).

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 producing 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 costliness 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 slag 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 Rachette. 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 extending the length whilst maintaining a relatively small width.

For some time development proceeded along these lines slowly and with much caution, chiefly 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, still undergoing development.

Fig. 34.—Development of the Blast Furnace (Gowland).

Here in 1902, blast furnaces 15 feet long by 56 inches wide were erected, the plant eventually consisting of seven such furnaces built in a straight line, and situated 21 feet apart from each other. A largely augmented ore supply subsequently coming to the smelter for treatment, an increased furnace capacity was required, for which only a very limited suitable space was available. Mr. E. P. Mathewson, 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 + 21 + 15, 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, 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 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 the working, and reduced the capacity.

The preliminary tests being satisfactory, the necessary constructional 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 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 joining up the other two 51-foot furnaces, so as to make a single one 123 feet in length, but certain difficulties in the matter of bringing coke supplies to the two sides, under the special conditions of available 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 subsequent furnace management or operation.

Figs. 35 and 36 indicate in plan and elevation the arrangement of the plant and accessories for these extended furnaces. Each 15-foot furnace had its own settler situated in front, and these have been retained without any change of position or any further additions. The hearth of the newly bridged portion slopes from the middle of the bridge to the tap-holes of the old furnaces, which still serve this purpose 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 side water-jackets of the old furnaces remain, being built up in two sets of panels, each 7 feet 6 inches wide, whilst the new bridge portions are constructed of three sets of jackets, each 7 feet wide.

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.

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 furnaces 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 11 per cent. was required formerly on a charge, only 10 per cent. was necessary under the new conditions.

(iv.) The rapidity of working of the furnace has increased owing to the effect of the narrow 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 are hotter than in smaller furnaces. In consequence more siliceous slags can be run, thus saving the cost of the fluxes which might otherwise be necessary.

(vi.) Marked decrease in incrustation. Crusting is most likely to occur at points where the smelting activity is lowest, and in the cooler parts of the furnaces, such conditions being usually prevalent at the corners, where the shape also assists in the holding up of material. Crusting is one of the chief troubles to be prevented and overcome in operating the blast furnace.

The elongated furnace of 87 feet length practically takes the place of five shorter ones, representing no less than 20 corners and 10 end jackets; 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 opportunity 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.