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By C.P. Proctor, J. T. Crawford, J.A. Younkins

Coal-cleaning plants using the launder process generally wash the fine coal (minus ? or minus 5/16-in.) separately in a plant consisting of washing launders or troughs placed one below another and sometimes supplemented by additional washing equipment such as tables and jigs. The coal and its associated impurities enter the feed end of the top launder, with a current of push water. As the material moves along the impurities, being heavier than coal, settle toward the bottom of the launder and are drawn off, with some coal, through "free discharge" boxes to the launder below. These boxes were given their name because their discharge falls freely; that is, it does not discharge into a water seal. Thus the impurities are progressively concentrated in the lower launders until box products are sufficiently free of coal to be discharged to refuse. The overflows of the upper launders go to clean coal while the overflow of the lowest launder, and perhaps that of the one immediately above, is usually recirculated to the plant feed. Thus the bulk of the clean coal is removed rapidly and the impurities are concentrated more slowly. The chief purpose of this paper is to present the cleaning data of such a plant under different operating conditions. OpeRating Equipment The washing units considered here are part of a complete plant cleaning minus 4-in. coal. The primary cleaned product is a high-quality gas and metallurgical coal and another portion of the output may be shipped as a steam coal. The units consist of two parallel Rheolaveur free-discharge units each containing five launders, together with three 7 by 14-ft. Plato tables. The launders are designated by the letters A-B-C-D-E (Fig. I), launder A being the top and longest; E being the bottom and shortest. Launder details are given in Table 1. The launders are equipped with a series of free-discharge Rheolaveur boxes, which are integral parts of the launder (Figs. 2 and 3). Each box with-

Jan 1, 1940

By C. P. Proctor, J. A. Younkins, J. T. Crawford

Coal-cleaning plants using the launder process generally wash the fine coal (minus ? or minus 5/16-in.) separately in a plant consisting of washing launders or troughs placed one below another and sometimes supplemented by additional washing equipment such as tables and jigs. The coal and its associated impurities enter the feed end of the top launder, with a current of push water. As the material moves along the impurities, being heavier than coal, settle toward the bottom of the launder and are drawn off, with some coal, through "free discharge" boxes to the launder below. These boxes were given their name because their discharge falls freely; that is, it does not discharge into a water seal. Thus the impurities are progressively concentrated in the lower launders until box products are sufficiently free of coal to be discharged to refuse. The overflows of the upper launders go to clean coal while the overflow of the lowest launder, and perhaps that of the one immediately above, is usually recirculated to the plant feed. Thus the bulk of the clean coal is removed rapidly and the impurities are concentrated more slowly. The chief purpose of this paper is to present the cleaning data of such a plant under different operating conditions. OpeRating Equipment The washing units considered here are part of a complete plant cleaning minus 4-in. coal. The primary cleaned product is a high-quality gas and metallurgical coal and another portion of the output may be shipped as a steam coal. The units consist of two parallel Rheolaveur free-discharge units each containing five launders, together with three 7 by 14-ft. Plato tables. The launders are designated by the letters A-B-C-D-E (Fig. I), launder A being the top and longest; E being the bottom and shortest. Launder details are given in Table 1. The launders are equipped with a series of free-discharge Rheolaveur boxes, which are integral parts of the launder (Figs. 2 and 3). Each box with-

Jan 1, 1940

By Harold A. Linke

The following article presents notes and data compiled and computed by the writer for use in the determination of: size and slope of mill launders, details of junction boxes and downspouts, and distribution of pulp. It also includes pulp formulas, volume of discharge through bushings, paths of discharging streams, banking of turns and curves. In the design of a concentrating mill, the draftsman usually considers launders to the extent of providing ample "head" for transmission of pulp from one department of the mill to another. The superintendent of construction aims to build his launders wide and deep, and on a slope steep enough to ensure positive flow. So launders are laid generally on too steep a slope, with the result that the repair gang works to maintain launder bottoms and sides against the erosive action of the conveyed pulp. To this end many kinds of liners have been devised: concrete, mastic, white iron, glass, rubber. True, the reduction of slope to the point where a pulp will flow with minimum wear, will minimize labor and expense; but when the desired slope is not predetermined, the millman hesitates to readjust his launder slopes by cut-and-try methods. To the end that the desired size and slope may be found readily, the writer has: (1) conducted experiments with pulps of different specific gravity, screen analysis and dilution; (2) gathered a few graphic data from standard reference works (such as Peele's Mining Engineer's Handbook and Taggart's Handbook of Ore Dressing); (3) computed several tables; and compiled results that have proved dependable for Utah Copper ore. It is quite possible that the following tables and graphs may not be applicable to all ores, but it is hoped that they may prove helpful, even in slight degree, in alleviating a portion of the trouble and migraine of some other millman. Launder Slopes The main object in the study of launder slopes is to provide grade steep enough to prevent "sanding" by a safe margin, and to avoid

Jan 1, 1943

By Harold A. Linke

The following article presents notes and data compiled and computed by the writer for use in the determination of: size and slope of mill launders, details of junction boxes and downspouts, and distribution of pulp. It also includes pulp formulas, volume of discharge through bushings, paths of discharging streams, banking of turns and curves. In the design of a concentrating mill, the draftsman usually considers launders to the extent of providing ample "head" for transmission of pulp from one department of the mill to another. The superintendent of construction aims to build his launders wide and deep, and on a slope steep enough to ensure positive flow. So launders are laid generally on too steep a slope, with the result that the repair gang works to maintain launder bottoms and sides against the erosive action of the conveyed pulp. To this end many kinds of liners have been devised: concrete, mastic, white iron, glass, rubber. True, the reduction of slope to the point where a pulp will flow with minimum wear, will minimize labor and expense; but when the desired slope is not predetermined, the millman hesitates to readjust his launder slopes by cut-and-try methods. To the end that the desired size and slope may be found readily, the writer has: (1) conducted experiments with pulps of different specific gravity, screen analysis and dilution; (2) gathered a few graphic data from standard reference works (such as Peele's Mining Engineer's Handbook and Taggart's Handbook of Ore Dressing); (3) computed several tables; and compiled results that have proved dependable for Utah Copper ore. It is quite possible that the following tables and graphs may not be applicable to all ores, but it is hoped that they may prove helpful, even in slight degree, in alleviating a portion of the trouble and migraine of some other millman. Launder Slopes The main object in the study of launder slopes is to provide grade steep enough to prevent "sanding" by a safe margin, and to avoid

Jan 1, 1943

By T. C. Campbell

Copper is extracted from the Chuquicamata oxide ore by a hydro-metallurgical process. The ore is crushed to 3/8-in. sizing, and leached with a sulfuric acid electrolyte. Chlorine is precipitated. and the ferric iron reduced in the enriched electrolyte, after which the copper is recovered by electrolysis with insoluble Chilex and lead-antimony anodes, the spent electrolyte being returned to leaching. Cathodes are melted and refined in market furnaces and cast into commercial wire-bar and cake shapes. Sulfuric acid for the process is supplied by the brochantite in the ore. Water for washing the ore is advanced through the solution system and finally after cutting by electrolysis to from 6 to 16.0 grams per liter is completely stripped by the cuprous chloride method and run to waste. The cuprous chloride so obtained plus that resulting from precipitation of chlorine from strong solution is dissolved in ferrous chloride brine and the copper cemented on scrap iron. The cement so obtained is in part used to reduce ferric iron in electrolyte, and as it is of exceptional purity is also furnace refined and cast into an exceptional quality of fire-refined copper. Fig. 1 is flow sheet of the operation. The reduction plant is composed of seven divisions as follows: (1) crushing, (2) leaching, (3) tailings disposal, (4) dechloridizing, (5) sulfur dioxide treatment, (6) electrolytic tank house, (7) smelting and melting. A brief description of each of these divisions will be given. Crushing Ore that is frequently blocky and up to 5 ft. in section is crushed in four stages to 3/8-in. mesh. The first stage of crushing is called primary and the next three stages are known as secondary crushing. The first stage is accomplished with two 60-in. Superior McCully gyratory breakers by which the ore is broken to 9 in. The second coarse gyratory breakers crush the ore to 3-in. sizing. There are seven No. 10 McCully gyratories in this service, for one of which recently has been substituted a No. 7 cone crusher. The third and fourth stages of crushing are accomplished by fifty 48-in. disk crushers. In the third stage 14 disk crushers break the ore to 1 in., and in the last stage 36 disk crushers turn out the final product of 0.371-in. sizing. Reference to the flow sheets of Figs. 2 and 3

Jan 1, 1934

By E. J. Duggan

It is a coincidence that the two plants in this country for treating copper ores by ammonia leaching came into existence at almost the same time, that of the Kennecott Copper Corporation at Kennecott, Alaska, in March, 1916, and that of the Calumet & Hecla Mining Co. in July of the same year. The success of both from their very beginning attests the merits of the process under widely different sets of conditions. As would be expected, in the time they have been operating these two pioneers in this field have greatly perfected their methods and improved their technic. The early difficulties inevitably encountered in undertakings of this nature have been largely surmounted and the process may now be considered to be well seasoned and matured. There have been no other commercial plants built in this country and but one elsewhere, the Bwana M'Kubwa in Northern Rhodesia, for the reason that this method of treatment may be employed on comparatively few ores. However, it has certain advantages, some of which are inherent to sand leaching processes generally and some to ammonia leaching in particular. Among these may be enumerated: (1) the avoidance of fine grinding; (2) the small amount of water required; (4) a high-grade product; (5) little likelihood of deleterious gangue constituents and fouling of solutions. It is true that great progress has been made in recent years in the flotation of the ores that come within the province of ammonia leaching. Nevertheless ammonia leaching has its possibilities and these may be overlooked in the present tendency toward the use of flotation for all ores. Probably the chief requisite to its further extension to suitable ores will be a cheap source of steam. Ammonia, the only reagent required, is unlimited in supply, readily transported and the price that is likely to obtain in the future should not make its cost at all onerous. Choice of Method for Kennecott Ores The Kennecott ores are typical limestone replacement deposits. The milling ores have a copper content of approximately 8 per cent in a limestone dolomite gangue. About half the copper occurs as sulfide, mostly chalcocite with some covellite, and the remainder in the form of the two carbonates, malachite and azurite. A comparatively coarse grind followed

Jan 1, 1934

By A. E. Wheeler, H. Y. Eagle

The copper industry in the Province of Katanga, in the Belgian Congo, which is now controlled and operated by the Union Minére du Haut Katanga, had its inception many years ago in the vision of a Scotsman, Robert Williams, now Sir Robert Williams, Baronet. He it was who personally negotiated the mineral concession with King Leopold of Belgium, sent in the prospecting parties who located most of the present mines of the company, later arranged for the Belgian participation from which the present company was formed, and has been its active mentor ever since. The idea back of his efforts was the establishment of an industry on the Congo-Zambezi divide, of sufficient importance to draw up to it the head of the Cape to Cairo Railroad, the dream of Sir Cecil Rhodes, and the consummation of which he had left as a legacy to Mr. Williams, his successor. The reports of the prospecting expeditions showed that copper in profitable quantity was there, but many difficulties lay in the way of the establishment of the infant industry. The isolated location of the mines, various tribal wars, the Boer War, and many other factors contributed to making progress very slow. However, in spite of these difficulties the railroad was finally built and a blast-furnace plant erected at Lubumbashi, near Elisabethville, the capital of the Province. Operations of this plant were started in 1911, treating high-grade oxidized ores and producing pig copper by a reducing smelting. Production has been continuous since that time and the company has grown so that at present it ranks as the largest individual copper producer in the industry. It is a fitting monument to the foresight, optimism and driving force of its founder, and the progressive policies of his Belgian colleagues. In January, 1914, the Union Miniere du Haut Katanga engaged as its consulting engineer, Archer E. Wheeler, metallurgical engineer, at that time general superintendent of the Great Falls plant of the Anaconda Copper Mining Co. Mr. Wheeler was commissioned by the Board of Directors to proceed into Katanga, and make all necessary investigations to enable him to report on a comprehensive development program, and to recommend a scheme of treatment particularly for their low grade

Jan 1, 1934

By Harold W. Aldrich, Walter G. Scott

The leaching process of the Inspiration Consolidated Copper Co. is based upon the solubility of oxidized copper in sulfuric acid and the solubility of sulfide copper, largely in the form of chalcocite, in ferric sulfate solution. This necessitates the carrying of a rather large amount of iron in solution. The ferric sulfate is regenerated from ferrous sulfate by the oxygen given off at the anode during electrolytic precipitation of the copper. In September, 1924, excavation was started at Inspiration, Arizona, for the construction of a plant to treat 7500 tons of ore per day. When the plant was begun, the fact developed that by a minor change in the height of the overflow on the leaching tanks, 9000 tons could be charged into each tank and treated satisfactorily. This determined the maximum capacity of the plant as 9000 tons per day. The first ore was charged in October, 1926, and there are now available the records of five complete years of operation; namely, 1927 to 1931, inclusive. The original estimate of copper content of the ore to be mined for this plant was 1.19 per cent, of which 0.77 per cent was as oxide and 0.42 per cent as sulfide. From the metallurgical results obtained at the 35-ton pilot plant, it was estimated that a 90.0 per cent extraction of the oxide copper and a 75.0 per cent extraction of the sulfide copper could be made. It was estimated that the tailing would carry 0.182 per cent total copper, resulting in a

Jan 1, 1934

By F. N. Koepel, L. A. Callaway

The property of the Andes Copper Mining Co. is in the central part of the Republic of Chile, Province of Atacama, about 75 miles to the eastward from the port of Chafñaral. At this point the Andean system is made up of two principal ranges, and the mine and metallurgical plant are both on the westerly slope of the outer or western range. Mean elevation of the mine is 10,500 ft., while at the plant the elevation ranges from 9445 to 9700 ft. A meter gage railroad extends from the company's port at Barquito, near Chañaral, to the reduction plant at Potrerillos. This road has a total length of 94 miles, of which 57.79 miles are owned by the Andes Copper Mining Co. and the remainder of the road is controlled by the State. Development of the Potrerillos mine and installation of the metallurgical equipment presented unusual difficulties due to the arid nature of the country, high altitude and remoteness of the location from any considerable center of population. As a primary step it was necessary that railroad transportation be provided in order that mine development and plant construction could go forward on an economical basis. Provision of an adequate supply of water for domestic and metallurgical uses involved the construction of approximately 178 miles of pipe lines, all of which have their sources at altitudes ranging from 12,000 to 14,000 ft. above sea level and traverse rough and difficult terrane. Water for metallurgical use is brought to Potrerillos from the Ola River, a stream originating in the high cordillera to the eastward from the property. It is delivered by means of a steel pipe line, 32 miles long, having diameters ranging from 22 to 36 in. Domestic water is supplied to the various camps through a system of pipe lines aggregating 146 miles in length, the Ola water being unfit for this purpose. These conductors gather water from numerous springs in the high mountains to the eastward from Potrerillos and deliver it to the mine, the reduction works and the port at Barquito. Extensive permanent housing facilities had to be provided as it was necessary to care for more than 3000 workmen and employees at the

Jan 1, 1934

By John D. Sullivan

Leaching of copper ores is a comparatively old art, probably dating back to medieval times. The leaching of mixed oxidized-sulfide ores, however, is modern. The first modern plant leaching mixed ores was at Ajo, Ariz., where commercial operations commenced in 1917. While the ore at Ajo was primarily oxidized there was always some sulfide present. The Inspiration leaching plant, which was put in operation in 1926, may be considered the first plant constructed for the extraction of both sulfide and oxidized copper. The pioneers in leaching mixed ores had a difficult task before them since they had very little fundamental data on the chemical and physical factors of leaching. For the very large amount of painstaking experimental work that was done before building the Ajo and Inspiration leaching plants the reader is referred to the original articles.' For a period of several years the writer worked on problems connected with the chemistry and physics of leaching mixed oxidized-sulfide ores. This paper summarizes the experimental work and discusses various factors involved in leaching ores. It also presents many data heretofore unpublished. Among the new data are those on chrysocolla, pages 518 to 519, inclusive; dioptase, page 519; covellite, page 529; virtually all those on chalcopyrite, pages 531 to 533, inclusive; enargite, page 533; and tetrahedrite and tennantite, pages 534 and 533. The oxidized copper minerals of commercial importance in American ores are azurite, Cu3(OH)2(CO3)2; malachite, Cu2(OH)2CO3; chrysocolla, CuSi03.2H2O; tenorite (or melaconite), CuO; and cuprite, Cu2O. Dioptase, CuSiO3.H2O is less common. Brochantite, CuSO4.3Cu(OH)2, while the chief mineral in the ore at Chuquicamata, is not of major importance in the United States. Metallic copper, while chemically not so, is frequently classed as an oxidized mineral of copper. The most important sulfide minerals are chalcocite, Cu2s; covellite, CuS; bornite, Cu5FeS4; and chalcopyrite, CuFeS2. Less important complex sulfides are enargite,

Jan 1, 1934

By Arthur S. Dwight

We author of this historical chapter will perhaps be criticized for limiting his subject to the history of the blast furnace in lead smelting, and especially to the part it played in the great custom ore-smelting business which for three decades or more was the means of opening up vast mineral treasures and practically dominated the mining industry in the Rocky Mountain region of the United States, in Mexico and in Canada. The men of that era shared the romance and the thrills of mining, and bore their part as pioneers in the building of the great West. The present writer was too young to participate in the first of the three decades mentioned, but immediately thereafter came under the apprenticeship of some of the men who had worked out the early metallurgical problems of the industry, became personally acquainted with most of the others of that group, heard the tales of early experiences, successes and failures, the gossip of the frontier and mining camps, from all of which there remain embedded in his mind clear impressions of the outstanding personalities of that time. That is the background against which the fifty years of intimate contact with the mining and smelting business and his own experiences as metallurgist stand out in his memory, and it is hoped that the gentle reader may understand and approve his desire to rescue from comparative oblivion the fundamentally valuable results accomplished by these fathers of the industry. It will be the purpose of this chapter first to narrate the general history of the growth of the industry; second, to pay tribute to some of the pioneer metallurgists, and third, to discuss some of the salient points of the metallurgical practice at different periods, to illustrate the technical progress of the art. The series of eight Annual Reports of Rossiter W. Raymondl,† U. S. Commissioner of Mining Statistics from 1869 to 1876, contain most of the available information about the early smelting operations in the Rocky Mountain region, and much is in the early volumes of the Transactions of the American Institute of Mining Engineers. The monograph by S. F. Emmons2 though not published until 1886, contained an elaborate description of the smelting works operating in 1879. Otto H. Hahn3 wrote a valuable paper in 1883, on smelting prac-

Jan 1, 1937

By G. U. Greene

The various lead producers have given the subject of lead drosses much attention in recent years but the problem of their economical treatment is yet to be solved. Formerly the copper in the furnace feed was removed from the furnace in the form of matte. The speiss at that time consisted principally of the iron arsenide. The practice today is to roast off most of the sulphur and the copper in the feed forms a dross in the furnace, which is removed from the furnace with the lead. This material is poured into the drossing kettles and allowed to cool down to but slightly above the freezing point of lead. The copper dross, containing almost all of the arsenic and part of the antimony, is skimmed off and treated in a small reverberatory furnace at most plants. This is essentially a liquation process. The skimmed dross contains from 70 to 80 per cent lead, and much of this lead is liquated out. The products from this treatment are: (1) impure lead, (2) speiss and (3) matte. The speiss from the reverberatory furnace was studied in this research. The literature on the subject is of very little assistance, as almost no work has been done on the problem of establishing the exact identity of the various phases present. One author(1)† states that speiss consists of a series of isomorphous compounds of the general formula FexAsy in which the Fe may be replaced by Ni, Co, Cu, Pb, Ag, Au, Pt, Zn, Cd and the As replaced by Sb, S and Bi. Hofman(2) states that the iron arsenide present is Fe5As. Experimental Work In the present research, the author used for methods in an attempt to determine the composition of the various phases: (1) chemical, (2) microscopic, (3) roentgenographic and (4) thermal. Previous to this research, considerable time had been spent making analyses of different speisses, but unfortunately no very definite conclusions could be made as a result of these analyses.

Jan 1, 1937

By T. E. Harper, Gustave Reinberg

Electrolytic refining of all lead bullion produced at the Oroya smelter of the Cerro de Pasco Copper Corporation has been adopted as standard practice. A pilot refinery with a capacity of 25 metric tons daily has been in operation since June, 1934, and a new refining unit having an initial capacity of 100 tons per day is under construction. This program has been made possible by certain modifications of the usual Betts refining process1, developed at Oroya, which have the effect of making the purity of the lead produced, and the acid loss incurred in operating the process, entirely independent of the amount of impurities in the original bullion. The modified process, as developed and at present used, treats impure bullion direct, without presoftening or other prerefining step, except that if the bullion carries tin it is desirable to detin before casting into anodes, since tin carries over to some extent to the cathodes. The refining process developed differs from usual Betts practice in two essential particulars: the anodes are cast thinner and replaced at more frequent intervals, and anode slimes are washed by a system of counter-current dipping, prior to their removal from the surface of the anode scrap. Limiting the period of anode corrosion insures a cathode of high purity, since solution of impurities at the anode can occur only if the thickness of the anode slime blanket, and hence the voltage drop across it, becomes greater than a certain critical value. The dip washing system operates to prevent acid losses, by recovering the highly concentrated, nearly neutral lead fluosilicate solution present in the pores of the slime blanket on the corroded anode, and doing this in such a way that there is no opportunity for chemical decomposition to occur. The best illustration of the efficacy of these two expedients may be secured from records of more than a year's operation of the 25-ton pilot plant, during which period the acid losses were less than 2 Ib. of 100 per cent hydrofluosilic acid (H2SiF6) per ton of lead produced, and typical assays of bullion and refined lead were as shown in Table 1.

Jan 1, 1937

By Galen H. Clevenger

Equipment for treating lead ores was added in 1905 to a copper plant which the United States Smelting Refining and Mining Co. had built in 1902 at Midvale, Utah, about 12 miles to the south of Salt Lake City. By 1906 the lead section of the plant had been rounded out and consisted of a 90-ton gravity concentrator, fifteen hand-rabbled reverbera-tory roasting furnaces, six blast furnaces, a briquetting plant for handling flue dust and fine ores, a regulus furnace for concentrating the low-grade matte made in the blast furnaces, and a baghouse. The copper section was later abandoned and the plant has been operated for many years exclusively as a lead smelter. Some 45 years had elapsed at the time this plant was built since the first attempts to smelt lead ores in the West, and lead smelting had become a large and profitable industry but fundamentally little that was really new had been developed, which could be embodied in a plant built in this period. Principles and tools to work with had remained much the same as those borrowed earlier from European practice. Capacities had been increased and there were mechanical improvements but hand charging was still in vogue and there was little attempt to recover suspended particles carried out with the furnace gases, beyond settling in flues, which usually was only partly effective for recovery of the coarser particles. Satisfactory roasting and preparation of the fine components of the charge were still partially solved problems. The Midvale plant of 1906 may be regarded as more or less typical of the smelting plant of this period for treating complex western lead ores as distinguished from the recovery of lead from the simple lead ores to the eastward. The operation of a concentrating mill directly in conjunction with a lead smelter was, however, not common practice. The ores treated in this plant have from the beginning come from mines owned and operated by the company or have been custom ores purchased in the territory tributary to the plant. The treatment of custom ores is an important service to the small producer who does not have sufficient tonnage to warrant a plant, and serves to increase the tonnage of the operator of large properties, thereby helping to reduce

Jan 1, 1937

By H. E. Lee, P. C. Feddersen

Blast-fuRnace operation at the Bunker Hill smelter deviates somewhat from common practice. Under existing conditions it is economical to operate the furnaces at "forced" capacity. The furnace feed consists chiefly of sinter and averages close to 50 per cent lead. Experience in smelting such material favors a low charge column attended by smaller proportions of coke and a low blast pressure. Under such conditions the gain in furnace tonnage greatly overbalances the somewhat higher slag losses and increased percentage of material fumed Furnace Charge The ingredients for the blast-furnace feed are stored in a group of 24 charge bins (3600 cu. ft. capacity). Each bin has two discharge chutes. Beneath each chute is a steel weighing hopper, suspended from a Fairbanks scale. The several hoppers are kept filled with the exact amount of material required to complete a furnace charge. The weighed material is discharged into a standard-gauge, bottom-dump charge car containing two compartments, each of which holds a complete charge (approximately 1½ tons). It is hauled to the blast furnace by a 44011 electric trolley locomotive. The proportionate parts of the various constituents of a representative blast-furnace charge are shown in Table 1. Table 1.—Representative Blast-furnace Charge at Bunker Hill Smelter Material Pounds Sinter.................................................. 2500 Limerock............................................... 0- 50 Coke................................................. 250-300 Scrap iron.............................................. 0-25 Settler skulls.......................................... 75 Undried zinc-plant residue................................ 0-200 Siliceous ore............................................ 25 Baghouse dust.......................................... 0-50 By-product slag......................................... 100 A variable stock tonnage of zinc-plant residue and return by-products results in fluctuations in the percentages of these materials in the blastfurnace feed. Accompanying such variations are corresponding increases or decreases in the amounts of coke, limerock and scrap iron necessary to obtain good furnace operation and adequate reduction.

Jan 1, 1937

By Jesse O. Betterton

In the Parkes process of lead refining, after desilverization has been completed by means of zinc additions, there will remain in the lead from 0.5 to 0.6 per cent zinc. At this stage in the refining practically all impurities are reduced to refined lead specifications except zinc and antimony. All excess of copper and nearly all the arsenic goes out with the silver. Silver remaining should not exceed 0.1 oz. per ton. It thus remains only to remove zinc and antimony to produce refined lead. The older practice was to complete refining by means of oxidation processes in reverberatory furnaces. Essentially this consisted of heating the lead to 1200' to 1300" I?. and allowing it to come in contact with a considerable excess of air passing through the furnace, or accomplish the same result by blowing the lead with steam by means of pipes introduced for that purpose, at the same time allowing air to enter the furnace. Usually some procedure involving alternate oxidation and reduction with fine coal or coke was employed in order to reduce the weight of dross or skim finally removed from the furnace. This practice had many elements of weakness in that the main flour of lead, after the preceding expensive operations had been performed, and the whole refining process nearly completed, was subjected to an operation at a comparatively high temperature, which involved much lead loss if the furnaces were not on bags, and the production of 3 to 5 per cent of the weight of the lead as dross or skim, which had to be resmelted and the lead either refined or produced as antimonial lead. In addition, the operating expense of the process was a considerable item. The process of dezincing with chlorine was developed as a substitute for the foregoing described oxidation process, and finally resulted in such advantages as nearly equal operating costs, a practical elimination of the lead loss, and the direct production of a very valuable byproduct, zinc chloride, that may be sold as such or retreated for its zinc and chlorine contents, both of which can be reused in the refining processes. Essentially this process consists of circulating the lead by means of a pump in a kettle through a closed reaction cylinder, gaseous chlorine entering the reaction cylinder with the stream of lead from the pump. Molten zinc chloride is thus produced and is discharged together with

Jan 1, 1937

By G. K. Williams

The continuous process of lead refining as at present operated at the Port Piric plant of The Broken Hill Associated Smelters Proprietary Ltd. was a development from investigations conducted in connection with the ordinary Parkes desilverizing process. It is not proposed here to discuss the theoretical considerations involved in continuous desilverizing because this aspect has already been dealt with in previous publications.1 Continuous Softening Process For economical desilverization of lead bullion, the bullion must first be subjected to refining processes for the removal of such impurities as copper, sulphur, arsenic, antimony and tin. This refining is effected by first drossing to remove copper and sulphur and then preferentially oxidizing the arsenic, antimony and tin from the bullion in the softening process. The theoretical basis of the continuous softening process is illustrated by the portion of the Pb-Sb-O ternary equilibrium diagram shown in Fig. 1. Under conditions of complete liquefaction, metal of the composition of softened bullion is in equilibrium with an oxide slag containing about 9 per cent antimony. In practice, of course, arsenic also is always present, and while the 0Pb-Sb-As-O quaternary system has not been worked out in detail, it has been found that in softening drossed smelter bullion of the composition produced at Port Pirie, arsenic is always reduced to 0.001 per cent or less when the antimony in the bullion is reduced to a satisfactory figure. In the operation of the continuous softening process, a two-phase system of a bullion low in arsenic and antimony and an oxide slag relatively high in those components exists in equilibrium. To this system is added continuously a bullion containing 0.2 per cent As and 0.8 per

Jan 1, 1937

By Jesse O. Betterton, Y. E. Lebedeff

AS a matter of the most widespread interest to lead-refining metallurgists, the process of desilverizing lead originated by Parkes is the most fundamental step in lead refining. While this basic operation, in addition to effecting a practically complete recovery of gold and silver, is also a great purifying operation and a governor on the purity and uniformity of the final refined lead, it has the important deficiency that no bismuth is removed in the operation. The problem of finding a satisfactory, inexpensive method for removal of bismuth that would fit in with the basic desilverization step has been one of enticing interest for lead-refinery metallurgists for many years. Of late it has become one of imperative interest, and while many attempts and much progress have been made, it is probably fair to say that metallurgical progress thus far is only catching up with commercial requirements. The Pattinson process, an original competitor of the Parkes process for desilverization, has received much attention for its possibilities for bismuth removal. Bismuth concentrations in the Pattinson process fall far short of the silver concentrations and as a result too many operations are required and a crawl in grade in the kettle units is experienced which requires periodic correction. Altogether, it has proved too expensive a remedy, from the cost standpoint of both installation and operation. The Betts electrolytic process is a fully satisfactory remedy but its present status in the field of lead refining is more a matter of bismuth removal than preference as a refining system, except, of course, in instances particularly favorable to the basic requirements of the process. The Betts process, of course, is essentially different from the Parkes process, and very difficult to associate therewith to any economy except in minor ways. Accordingly, where economy dictates the Parkes process, and this is true of far the greater portion of the world's tonnage, the adoption of the Betts process primarily for bismuth removal is apt to prove an expensive remedy on account of its normally higher cost of operation, inelasticity of the equipment for variable tonnages, and high per ton cost of installation.

Jan 1, 1937

By G. L. Oldright, T. B. Brighton, C. M. Dice

The base bullion from the lead blast furnace contains varying amounts of the impurities left in the smelter feed by the concentrator, regardless of the method of smelter operation. These impurities may be held in part in mechanical suspension rather than being present as true alloys. However, usually there is little free space in the crucible in which to separate these drosses before the bullion is taken from the lead well. When such a dross is melted in a laboratory crucible and allowed to cool, the products obtained are lead bullion, speiss, and matte in well separated layers. Typical analyses of layers are: Analyses. Per Cent Element Bullion Speiss Matte Pb.............................................. 92.0 41.2 38.3 Cu.............................................. 2.4 36.3 40.6 As............................................... 4.0 16.5 4.3 Sb............................................... 0.7 0.2 0.4 S................................................ 0.6 1.9 13.1 Fe............................................... 0.0 0.0 0.0 Totals......................................... 99.7 96.1 96.7 The absence of iron in these products drossed from the blast-furnace bullion proves that the impurities for the greater part are not simply finely divided blast-furnace speiss or matte that has been dissolved or dispersed in the lead, but are held in solution by the liquid metal and rise to its surface as infusible compounds when the temperature of the bath is lowered. In general, the drosses made locally are metal sulphides and arsenides and not oxidized compounds. The large content of metallic lead also does not seem to be retained because of oxidized compounds

Jan 1, 1937

By P. F. McIntyre

Electrolytic refining of lead was first carried out on a commercial scale at Trail, B. C., in 1903. The capacity at that time was small, and cell-changing operations were managed with chain block and monorail. Development up to 1922, utilizing crane installations and cell additions, attained a production of 100 tons per day. Successive additions from 1922 to 1926 further increased the rated capacity to 396 tons per day. Other improvements in metallurgical practice have gradually increased production, until for the year 1935 it was 448 tons per day. During the first five months of 1936, an average daily production of 504 tons was maintained without increasing the size of the plant. Lead Refinery Buildings.—The tank room and the lead-melting and casting equipment are housed in two buildings, which parallel each other and which for the greater part of their lengths are joined by a connecting aisle. All basements, walls, sumps and floors are of concrete. Where exposed to the electrolyte, the concrete has a protective coating of asphalt (220" melting point) over a coat of asphalt paint as a primer. Buildings erected before 1924 are of wood. More recent additions are made of structural steel with gunite siding. A brick building, adjoining the older section of the tank room, houses the slimes-treatment plant. The power house, containing motor generators, air compressors and rectifiers, is constructed of tile. Cell Construction.—The electrolytic cells are of reinforced concrete lined with asphalt. They are cast in pairs, the outside dimensions of a double cell being 113 in. long by 73 in. wide by 48 in. deep. The bottom is 6 in. thick and the sides, ends and center wall are 4 in. thick. A ledge 3 in. wide and 2½ in. deep is left along the top of the side walls to carry the insulators and bus bars. Reinforcing consists of ½-in. round iron rods. Cell inspection and repair work is done in conjunction with the cell-changing operations. The repair work necessary is usually confined to renewing chipped asphalt, unless a tank has been accidentally broken. Minor concrete repairs can be easily made in place. The broken concrete

Jan 1, 1937