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By W. B. Boggs, J. N. Anderson

An account of the operation of the Noranda smelter covering the period from December, 1927, when the smelter started, until December, 1929, was given in a paper published by the Canadian Institute of Mining and Metallurgy early in 1930.' The object of the present paper is to review matters of interest in the original paper and to describe the present construction and operation of the smelter, with particular reference to the reverberatory-furnace practice. The Noranda smelter was built by Noranda Mines, Ltd., from the designs of A. E. Wheeler, consulting metallurgist of New York, for the reduction of ores and concentrates from the Horne mine and to treat, on a custom basis, ores and concentrates from other mines in the Rouyn district. The smelter is at Noranda, about 330 miles directly north of Toronto. Construction work was started in June, 1926, and copper was first produced in December, 1927. The smelter was built throughout as two duplicate units, each with a rated capacity of 500 tons of ore per day. Operations were started with one unit and after several months it was found that one unit alone could smelt 1000 tons of ore a day. At that time there was only one hoisting shaft, No. 3, and its hoisting capacity and the capacity of the original crushing plant were insufficient to supply ore for the two units at the increased tonnage. Accordingly No. 4 shaft was sunk and a new crushing plant provided to handle the ore hoisted there. Operating two units at capacity also required the installation of two more converters to treat the additional quantity of matte produced. This work was completed in the latter part of November, 1929, and since then both smelter units have been operating continuously. In 1928, a mill to treat 500 tons a day of concentrating ore was completed. Since then the concentrator capacity has been increased twice, and at present 2000 tons a day of ore are being concentrated. The layout of the surface plant of the mine, concentrator and smelter is shown in Fig. 1 and the general plan in Fig. 2. The smelter is shown in more detail in Fig. 3. The original hoisting shaft, No. 3, is now used only for handling men, supplies and waste rock, and the original crushing

Jan 1, 1934

Data for 1962 world-wide, non-communist bloc geophysical activity, including geochemistry, are now available through the Society of Exploration Geophysicists. This information was made available to the AIME Geophysics Unit Committee through the courtesy of Charles L. Elliot, Mining Coordinator for the SEG. The data were compiled from 245 responses to 496 questionnaires mailed (49%). The data for the U.S. are a realistic representation of actual activity, followed by Asia, Canada, Europe, Australia, Africa and Latin America in order of decreasing reliability.

Jan 9, 1963

By Freeman Bishop

Nothing has disturbed the copper- lead-zinc producers in recent years more than smelter tolerances set by some states. What hurt most, copper spokesmen told the Government, was the Government's own erroneous estimates of smelter pollution controls. Government charts indicated that 23 firms smelting copper, lead and zinc may need to invest about $108 million by fiscal 1976 to control stack emissions. The copper industry's independent study, however, indicates that the required controls could cost as much as $550 million to install and $200 million a year to operate.

Jan 1, 1971

By R. Hoffmann

The Waelz process produces oxides of volatilizable metals from ores, metalliferous products and residues. The process was originally used for recovering zinc and lead, where tailings and residues containing those metals had accumulated at mines or smelters, but it is applicable for obtaining tin, arsenic or other volatilizable metals. It is useful when other methods of obtaining the metals from the ores or residues are too difficult or expensive or show too low a recovery. The Waelz process was developed in 1923 for the treatment of large calamine dumps in Upper Silesia, which showed zinc contents of 8 to 10 per cent. and lead contents up to 3 per cent. The rapid development of the process, from the first trials up to the plants now working or in course of erection, is shown in Fig. 1. The first industrial Waelz plant was started in 1925. Within two years the capacity of the working plants reached 700 tons material per day of 24 hr. During the year 1928 the daily capacity of the plants now at work or in course of erection will rise

Jan 1, 1928

By Harrison Everett Ashley

Slimes are usually defined as all material passing a certain sized sieve, which is invariably the finest sieve employed by each metallurgist in his tests; 100-mesh and 200-mesh have been taken as the limits by different writers in recent volumes of the Transactions; 150-mesh is common practice. The reader of the first two volumes of Richards' is impressed by the inflexibility with which all grinding-arrangements are praised or condemned according as they make little or much slime. Until recently, one of the principal aims of the mill-man was to keep down the slimes. He was obliged to have separate apparatus for sand- and slime-treatment. The latter was sometimes puzzling and erratic. Often it was found most economical to omit treatment of the slimes entirely, or to store them in immense basins or heaps, waiting for possible future improvements in methods of treatment. Recently, improved processes of agitation and filtration have made possible the successful extraction of values from many slimes. The tendency, when the metal is exceedingly finely disseminated, is to cut out all sand-treatment and to slime everything. These processes have been almost entirely mechanical in their conception, and it is the object of the present paper to show that certain chemical considerations may greatly hinder or assist their successful operation. It has long been a source of dissatisfaction to chemists that their analyses fail to be of much help to the clay-worker; that it is a matter of common occurrence for clays of closely identical ultimate chemjcal constitution to show the widest possible divergence in physical properties.

Jan 1, 1911

By J. Ralph Finlay

In latitude 4" S., about fifty miles from the Pacific, and in the amphitheater on the west flank of the Andes, which is drained by the river Tumbez, is the ancient town and mining district of Zaruma. The town is picturesque in the extreme, perched on a ridge 1800 feet high, between two lovely mountain rivers, looking out upon neighboring hills and valleys clothed in rich green, relieved by many hamlets roofed in red tiles, and, at greater distances on all sides, upon mountain slopes of gloomy forest crowned by the bare summits of the Andes. General Description. The gold-mines, discovered in 1549, mere worked extensively by the Spaniards in the 16th and 17th centuries, until their success was curtailed by the exhaustion of the oxidized surfaceores, and the uncovering of the refractory and hard ores underneath. The industry slowly declined until it was nearly or quite at a standstill in 1878, when an unfortunate English company, organized by Chilian promoters, acquired some of the best properties and began to work them. This company, after an extravagant début, relapsed into a struggle against poverty and mismanagement, and finally collapsed in 1894. In 1896 the English mines and many others, covering most of the district, were acquired by the South American co., formerly the South American Development Co., of New York. (See Figs. 1 and 2.) The vein-formation of Zaruma is extensive and complex, and the ores are somewhat refractory. In the immediate neighborhood of the gold-bearing veins the formation seems to be altogether volcanic, consisting of rather fine-grained holocrystalline diorite; but a few miles away, on either side, some gueisses and crystalline schists are to be found. About four miles south of the mines, and beyond some intervening granites, appear

Jan 1, 1901

By Frank F. Poland

TEE purpose of this paper-is to describe a new process for the refining of secondary copper-base metals and a specially designed high-temperature electric-resistor furnace used in the process for the distillation of zinc from these metals. The process was used for the conversion of nickel silver to cupronickel and to recover the zinc content in metallic form. It also proved applicable and economical for the treatment of scrap brass for separating and reclaiming the copper and zinc. The relatively large furnace developed for the distillation of zinc is described in detail, for it made economically possible the direct recovery of metallic zinc from scrap brass, which is a new step in the processing of such material. Considerable interest has been shown in the utilization of the distillation furnace for other metallurgical processes, where it is of advantage to operate in a controlled neutral or reducing atmosphere, or to avoid having the processed metal in contact with refractories other than carbon, or to operate at temperatures between 3000 and 4000°F. PURPOSE OF DEVELOPMENT It was recognized early in the war effort that nickel would become a critical metal, therefore all nickel stock and alloys containing relatively high percentages of nickel were frozen immediately. Among these alloys were considerable quantities of various grades of nickel silver, an alloy containing copper, nickel, zinc, and in some grades lead, also minor amounts of iron, manganese and other constituents. At the time there was no existing plant or process to eliminate the zinc, lead, and minor constituents so as to convert this metal to cupronickel, one of the important alloys needed by the U. S. Navy for condenser parts. Therefore the Bureau of Industrial Conservation of the War Production Board requested the executives of the leading nonferrous metal companies to have their metallurgical staffs study the problem of converting this metal to cupronickel conforming to U. S. Navy Specification 44T40. The composition specified therein is as follows: nickel, 29 to 32 per cent; zinc, 1.00; iron, 0.00 to 0.60; lead, 0.05; tin, 1.20; manganese, 0.00 to 1.00; total of other elements (maximum), 0.50; copper, remainder. The problem in converting these alloys to cupronickel was the elimination of zinc, lead, iron, manganese, and other elements without appreciable loss of nickel to slags or other by-products. Revere Copper and Brass Incorporated, through its Research and Development Department, undertook to solve this problem. A Government-owned plant was built at Rome, N. Y., and there the con- version was successfully made.

Jan 1, 1946

By Hugh Neill, O&apos

In previous papers,' the author has reported the results of experiments on the straining in tension of a single-crystal test piece, about 0.6 in. long, of vacuum-melted electrolytic iron containing 1.8 per cent. of silicon. The crystal was originally in the "as cast" condition, and was stressed slowly along a tetragonal axis so that slipping along conjugate gliding planes finally transformed the original cylindrical portion into a pair of fracture wedges. One of these fracture wedges was sectioned along the stress axis in a plane at right angles to the edge of the wedge. A metallographic study of this plane of section led to the conclusion that slip had taken place in the (111) direction along (112) planes making a V shape with the stress axis. The present paper describes the results of scratch-hardness tests conducted on the same section, in order to determine the distribution of strain hardness therein, followed by annealing, and considers the resulting variations in hardness. These observations supplement those made by Krivobok2 on similar material reduced 25 per cent. by hammering (i. e., by impact) in the cold. In order to investigate ageing effects, two small crystals of the same alloy were also tested. The Specimens Employed The principal specimen, A, was prepared for the metallographic examination leading to the determination of the slip planes of the crystal. Therefore the work described in this paper was conducted on metal which had been repeatedly etched and had been allowed to rest for about 30 months after straining. It consisted of a section of the fracture wedge and contiguous unstrained head of the test piece taken along a tetragonal axis parallel to the direction of stressing, the plane of section originally being not far removed from the (011). Several subsequent grindings

By Robert B. Kistler, Ward C. Smith

The borate industry is one of the few sectors of the mining and mineral-processing industry which the United States still dominates. Since about 1927, the United States has supplied over half of the world's total boron and borate raw materials, and many of the chemical derivatives as well. Chiefly responsible for this record are companies producing sodium borates from two large deposits in southern California, at Boron (Kramer) and Searles Lake. The sodium borates, or products derived from borax and borax-containing brines, are the preferred borates for industry because they are soluble in water and combine readily with most other chemicals. Calcium borates are required for certain end uses, and these are produced from smaller colemanite-ulexite-probertite bodies. Sodium-free borate compounds also are produced from boric acid, which can be manufactured from either sodium or nonsodium borate minerals. During the past hundred years, the use of boron compounds throughout the world has risen from a few tons per year to the equivalent of over a million tons of B2O3. Leadership in supplying borates shifted from country to country and from region to region during this period, generally as the result of the development of newly discovered deposits that were better in quality, more accessible, or more economically exploitable than previous sources (Table 1). Primary Products and End Uses The principal borate products offered by primary producers in the United States are listed in Table 2. In general, these products are utilized worldwide by almost all consumers, regardless of the original source of the primary borates, In the United States, the primary borates all originate in southern California. Colemanite is produced as crude or processed crude ore from the Death Valley area and is utilized by the fiber glass and ceramic industries without further treatment. Production from Searles Lake, CA is from brines that are processed in chemical plants where borax and other compounds are recovered and the principal borate products are produced. At Boron, CA, the world's leading producer is mining a deposit of mineral borax and operating a chemical refinery. The mineral borax is mined together with included clay, and refined borax or "10-mol" is one of the refinery products. The mineral and the chemical compound are identical, Na2B4O7 • 10H2O. In addition to ship¬ping all major refined borate products from [ ]

Jan 1, 1983

By Arthur Baker III

The orebodies are long bedding-plane lenses of chalcopyrite and sphalerite, associated with garnetite masses. Most of the orebodies are within a 50-ft thickness of Cambrian limestone; other Paleozoic limestones and dolomites are locally metamorphosed but only slightly mineralized. Pre-mineral faults are numerous, but shallow folds were the main ore-localizing structures. JOHNSON camp is in the northwestern part of Cochise County, Ariz., about 50 miles east of Tucson. The nearest major mining districts are Tombstone and Bisbee, respectively 27 and 50 miles to the south, and the Superior-Miami-Globe-Ray porphyry copper group, about 90 miles northwest. Like many mining camps of the Southwest, Johnson camp is said to have been worked by the Spaniards. The first production on record was in the early 1880's, when an unknown amount of oxidized copper-silver ore in the Peabody mine was mined from replacement orebodies in the Pennsyl-vanian Naco formation. From 1904 to 1911 outcropping oxidized ores in the Cambrian Abrigo formation were worked, and an estimated 100,000 tons of copper ore were shipped. In 1912 the first large sulphide orebody of the district, the Republic Manto orebody, was discovered, and in the following few years some 250,000 tons of predominantly sulphide ore were shipped. The average grade of this ore was approximately 4.5 pct Cu, 6 pct Zn, 0.8 oz. Ag, and 0.001 oz Au. The Republic, Copper Chief, and Mammoth mines were the principal producers during this period. Mining ceased in the district after 1920, and until 1943 only small-scale leasing operations were carried on. In 1943 all the mines that had been productive were acquired by the Coronado Copper and Zinc Co., the present operators, and in the 10 years since that time the district production has amounted to approximately 350,000 tons of milling ore averaging 2 pct Cu and 6 pct Zn. Most of this ore was produced from the Republic and Mammoth mines, but since 1950 a large part of the production has been from the new Moore mine. The total known production from the district, then, is about 3/4 million tons of copper and copper-zinc ore of low grade. All of this ore was produced from orebodies associated with garnetite in the middle member of the Cambrian Abrigo formation. In addition to this known production, an unknown tonnage of ore was extracted from the Peabody mine orebodies that lie in the Pennsylvanian Naco formation. Published information on the geology of the district is limited. Aside from brief references in various mining journals, only three papers on the district have been published.1-3 One of these is a U. S. Bureau of Mines report on a diamond drilling program, one is a brief paper on the general geology of the district, and the third is a report on geochemical experiments, with a section on the occurrence of the orebodies. The last two are by John Cooper, of the United States Geological Survey, who has done much detailed work in the area. Stratigraphy The rocks of the mineralized area are Paleozoic sediments ranging in age from Cambrian to Penn-sylvanian. Several disconformities are present in the stratigraphic column, the most important one being between the Cambrian Abrigo formation and the Devonian Martin formation. There are no angular unconformities. Within the district, the Paleozoic sediments lie in a fairly uniform monocline, striking northwest and dipping 30" to 50" northeast. This local monocline is part of a domal struc-ture centered in the Little Dragoon mountains to the southwest. The Texas Canyon stock, a quartz monzonite body intruded probably during the Laramide revolution, lies south of the mineralized area (Ref. 2, p. 33). The Paleozoic rocks dip away from the stock, and on the surface are separated from it by at least 1500 ft of Pre-Cambrian rocks. The outcrop pattern of the northeastern edge of the stock suggests that it may dip gently northeastward, passing below the mineralized area at moderate depth. No quartz monzonite has been found in mine workings or diamond drill holes, which reach to depths of 1000 ft. The only igneous rock found in the mineralized area is a lamprophyre dike cutting the Naco limestone in and near workings of the Peabody mine. With the exception of the lowermost beds—the Bolsa quartzite and the shaly lower member of the Abrigo formation—the Paleozoic sediments are predominantly carbonate rocks, Fig. 1. The middle member of the Abrigo formation, which contains the principal ore-bearing beds, is limestone, with thin shale partings throughout most of its 250-ft thick-ness. Near the top of this member is a sandy bed some 25 ft thick. The upper member of the Abrigo formation and the lower half of the overlying Devonian Martin formation are dolomitic, with num-

Jan 1, 1954

By L. D. Godshall

The latest revised edition of Mr. C. A. stetefeldt's book on the Lixiviation of Silver-Ores, which appeared very recently, contains no mention of the volatilization of silver in chloridizing-roasting—an omission which is the more remarkable in view of the fact that in former editions of the work this important subject was noticed. Moreover, Mr. Stetefeldt has discussed it to some extent in his paper on " The Stetefeldt Furnace," * in which he criticised certain statements made by me before the Colorado Scientific Society in May, 1893, in a paper entitled " A Review of the Russell Process;" ? and Mr. Morse, in his paper on " The Lixiviation of Silver-Ores by the Russell Process at Aspen, Colorado," $ has given statistics showing the loss of silver by volatilization (including the dustloss, which Mr. Morse estimates at not more than 1 per cent.) on more than 30,000 tons of ore roasted, to have been 9.16 per cent. The matter must certainly be deemed important enough to warrant me in returning to it, and offering, with the aid of the figures made public since my first paper, a reply to the criticisms of Mr. Stetefeldt. The passages of my paper which he deemed to be " ques-

Jan 1, 1897

By Christopher G. Atwater

The object of this paper is to describe and discuss the progress that has been made, up to the present date, in the development of the modern by-product coke-oven. There are few members of the Institute who have not heard more or less of the by-product oven-installation, and there are many whose acquaintance with it is intimate and practical. Since the general opinion on the subject has hardly become crystallized as yet, it is hoped that this paper will at least furnish matter for discussion. In order to obtain an idea of the present status of the industry in the principal iron-producing countries, the following table has been prepared : Table Showing Proportion of Coke Made in By-Product Ovens in Various Countries. United Kingdom (1898, Mr. George Beilby). Short Tons. Per Cent. Coal coked in by-product ovens,. .. 1,400,000 7.9 Coal coked in blast-furnaces with by-product recovery,.......2,240,000 12.7 Coal coked in non-by-product ovens,. . 14,000,000 79.4

Jan 1, 1903

By S. W. Ross

In 1936 a description of the plant (Fig 1) and process employed by the Electrolytic Zinc Co. of Australasia Ltd. for the recovery of zinc from zinc concentrate by the electrolytic process was prepared. † During the twelve years which have elapsed since the preparation of the earlier paper, several major changes in the metallurgy of the process have been introduced. It is the purpose of the present paper to give a general description of these changes and thus to bring up to date the description of the plant and process. Summary The major changes in Risdon practice since 1936 have been: 1. Replacement of two stage roasting by a preliminary roast followed by the flotation of all the leach residue and the roasting of the flotation concentrate. 2. Screening of all calcine fed to the pachucas. 3. Continuous leaching of calcine and improved classification of pachuca discharge. 4. Close control of hydrogen ion concentration during purification for iron removal. 5. Recovery of cobalt as a good grade oxide. 6. Production of part of the zinc output in the form of "four nines" metal (99.99 pct purity). 7. Closer spacing of electrodes thus increasing the potential output of cathode zinc per cell by 50 pct. Changes which are in prospect and for which construction work is proceeding at the present time involve the starting up of: 1. Two suspension roasters. 2. A contact acid plant to produce 150 tons‡ of acid per day and replacing the existing Mills Packard chamber plant. 3. Extra power station capacity to permit greater current flow to existing cell room units. This will increase the output of cathode zinc from about 245 to 290 tons per day. Plans for the future envisage the building of an ammonium sulphate plant, the first unit of which will produce about 50,000 tons per year, and improved treatment of zinc plant residue for the recovery of zinc, lead and other metals. At the end of this paper tables of metallurgical data are presented relating to the year ending June 30, 1948. Details of Changed Practice Output In 1936 the production of cathode zinc amounted to about 200 tons per day. This has since been increased to about 245 tons per day while plant extensions are practically complete which will permit of an output of about 290 tons per day in the near future. Roasting Division ROASTING POLICY A major change has occurred in the roasting policy. Twelve years ago the method in use was to carry out a two stage roast in the first stage of which sulphide sulphur was reduced to about 6 pct. The pre-roast calcine was re-roasted in modified Leggo furnaces using coal as fuel, sulphide sulphur being reduced to about 0.8 pct. The whole procedure was described on pp. 482491 of the earlier paper. Although this roasting procedure had certain advantages it possessed some distinct disadvantages. For instance, it appeared uneconomical to heat up the entire input of pre-roast calcine to roasting temperature by the expenditure of fuel in order to oxidise a few percent of sulphide sulphur. It was argued that if the pre-roast calcine were leached and a process could be developed for the recovery of a zinc sulphide concentrate from the leach residue, this concentrate, small in weight compared with the pre-roost calcine, would probably roast autogenously, thus virtually eliminating the expenditure of fuel as well as greatly increasing the weight of sulphur oxidised per square foot of furnace hearth area. The obvious method of producing a suitable concentrate from leach residue was by flotation. It will be recalled (see p. 495 of the earlier paper) that when two-stage roasting was practised the leach residue was classified, the granular fraction was ground and floated while the slime fraction was thickened, filtered and dried ready for shipment to a lead smelter. This process worked quite successfully. However, when trials were made of leaching a calcine carrying several percent of sulphide sulphur the granular fraction still floated well, but the slime fraction carrying 8-10 pct sulphide sulphur yielded very poor results when subjected to flotation. This fact held up the application of "pre-roast" leaching for many years. However, successful flotation of the slime fraction of leach residue was finally achieved and in August 1940 the slime flotation plant began operation, while the leach-

Jan 1, 1950

The 140th meeting* of the American Institute of Mining and Metallurgical Engineers was held in New York, Feb. 16 to 19, 1931. It consisted of the annual business session, twenty-six technical sessions at which 149 papers were presented, two round table discussions, two meetings of the Board of Directors, three meetings of Section Delegates, two formal lectures, four group dinners and 11 group luncheons, luncheon daily for members and guests, 20 committee meetings and conferences, the annual reception and dinner dance, an informal dance, a smoker-dinner, two meetings of the Woman's Auxiliary and a special program of entertainment for the ladies. Former members of the Rocky Mountain Club and friends lunched together at the Engineers Club on Tuesday. The proceedings of the Iron and Steel Division included four general technical sessions, an Iron Ore Round Table, the Howe Memorial Lecture, the Division luncheon at the Engineers Club Wednesday noon, a session each of the Executive Committee and the Committee on Nondestructive Testing, a luncheon each of the Committee on the Manufacture of Alloy Steel and the Blast-furnace Conference Group, and a dinner of the Committee on Physical Chemistry of Steelmaking. Dr. F. F. Lucas delivered the Howe Lecture, the title of which was "On the Art of Metallography." The Institute of Metals Division held three general sessions, its annual lecture, which was delivered by Dr. Arne Westgren and entitled "X-ray Determination of Alloy Equilibrium Diagrams," and a Division dinner at the Hotel New Yorker Thursday evening, at which Dr. G. W. Thompson was the principal speaker. The Executive Committee assembled at a luncheon on Thursday. A round table on Coal Land Valuations, a general technical session and a Division luncheon were held by the Coal Division. C. E. Bockus, President of the National Coal Association, made an informal talk at the luncheon which was held at the Engineers Club Monday noon. The Petroleum Division held six general sessions, including its annual production symposium. The annual dinner was held at the Engineers Club Thursday evening. A Gold Supply Symposium was the feature of the proceedings of the Committee on Mining Geology. This Committee also held a general session and a luncheon meeting. Sessions were held by other Technical Committees as follows: Mining Methods, 2; Mine Ventilation, 1 (besides which the Committee on Mine Ventilation Code held a meeting); Ground Movement and Subsidence, 1; Milling Methods, 1; Nonferrous Metallurgy, 1; Engineering Education, 1 (also a Committee luncheon); Non-metallic Minerals, 1, and Geophysical Prospecting, 2.

Jan 1, 1931

By Cyril Stanley Smith

For many years sulphur and its congeners selenium and tellurium have been regarded as very harmful elements in copper and copper-rich alloys. Like many similar prejudices, this has been found to be partly wrong. All three of these elements actually have no great effect on the working properties of copper and its alloys, and are desirable additions in that they considerably enhance the machinability of alloys in which they are present. Because of their possible commercial importance, a study of the machinability and other properties of the pure binary alloys was made, as described in this paper, and the effect of selenium on a large number of copper alloys was determined. The current opinion regarding the deleterious effects of elements in this group undoubtedly arose because of occasional failures of tough-pitch copper containing unusually high tellurium, and because of the well-known fact that sulphur causes spewing and unsoundness in tough-pitch copper. Siebe,1 also Deretchey and Margolina,2 studied the effect of sulphur on copper and showed that in the absence of oxygen it was present in the form of copper sulphide, which was less harmful to the properties than an equivalent amount of copper oxide. No data have been published on selenium and tellurium, although the statement is often made, following Eggleston,3 that tellurium makes copper hot-short. Copper specifications call for only traces of these elements. Skowronski and Mosher4 give data on the effect of selenium and tellurium on the conductivity of tough-pitch copper. In private communication Mr. Skowronski says that the samples were hot-forged from small castings and drawn to 0.081 in. diameter, but it should not be inferred from this that these elements are harmless to the working properties of wirebars and commercial shapes. Constitution and Microstructure Constitution.—Sulphur, selenium and tellurium exist in their alloys with copper in the form of copper sulphide, selenide or telluride. These compounds are practically insoluble in the solid state and possess limited

Jan 1, 1938

By Cyril Stanley Smith

For many years sulphur and its congeners selenium and tellurium have been regarded as very harmful elements in copper and copper-rich alloys. Like many similar prejudices, this has been found to be partly wrong. All three of these elements actually have no great effect on the working properties of copper and its alloys, and are desirable additions in that they considerably enhance the machinability of alloys in which they are present. Because of their possible commercial importance, a study of the machinability and other properties of the pure binary alloys was made, as described in this paper, and the effect of selenium on a large number of copper alloys was determined. The current opinion regarding the deleterious effects of elements in this group undoubtedly arose because of occasional failures of tough-pitch copper containing unusually high tellurium, and because of the well-known fact that sulphur causes spewing and unsoundness in tough-pitch copper. Siebe,1 also Deretchey and Margolina,2 studied the effect of sulphur on copper and showed that in the absence of oxygen it was present in the form of copper sulphide, which was less harmful to the properties than an equivalent amount of copper oxide. No data have been published on selenium and tellurium, although the statement is often made, following Eggleston,3 that tellurium makes copper hot-short. Copper specifications call for only traces of these elements. Skowronski and Mosher4 give data on the effect of selenium and tellurium on the conductivity of tough-pitch copper. In private communication Mr. Skowronski says that the samples were hot-forged from small castings and drawn to 0.081 in. diameter, but it should not be inferred from this that these elements are harmless to the working properties of wirebars and commercial shapes. Constitution and Microstructure Constitution.—Sulphur, selenium and tellurium exist in their alloys with copper in the form of copper sulphide, selenide or telluride. These compounds are practically insoluble in the solid state and possess limited

Jan 1, 1938

By C. M. Young

Bituminous coal piled in heaps or bins frequently undergoes a process of spontaneous heating as the result of the absorption of oxygen. It seems probable that the first absorption of oxygen by coal which has not previously been exposed to the air may occur as a condensation or a combination of oxygen in some form which does not result in the production of carbon dioxide, but slow combustion soon begins. The absorption of oxygen is accompanied by an increase of temperature, and this by an increased rapidity of absorption; hence the dangerous condition proceeds from bad to worse, until the kindling point is reached, unless the process is interrupted. A dangerous rise of temperature can be prevented by excluding oxygen, by increasing the bulk of coal in proportion to its surface exposed, or by circulating enough air to dissipate the heat produced. The storage of coal, in practice, varies from almost complete exclusion of oxygen, by storing under water, to such freedom of access as exists when the coal is stored in open piles. Oxygen available for absorption by coal is supplied by the air in the interstices between lumps and by additional air which may enter the pile through circulation. The size of the coal largely affects both of these supplies; for if the fragments are small the spaces between them constitute a small percentage of the total volume, and little oxygen will be available unless the circulation of air brings in a fresh supply, while with large fragments the percentage is relatively large; circulation also is much easier through a pile of coarse lumps. In the case of lump coal, although a large amount of air may be present, the exposed surface is comparatively small and there is little opportunity for the absorption of oxygen to be so rapid as to cause dangerous heating. Attempts have been made to prevent heating by allowing a sufficient circulation of air to carry off the heat generated; obviously this method cannot be applied to fine sizes nor mixed sizes. When storing coal at the University of Illinois, an attempt has been made to use so much fine coal and to pack it so thoroughly as to prevent circulation of air, while also reducing the original air in the pile so as to prevent heating. This method has been fairly successful, but in some

Jan 1, 1918

By Rudolf Gahl

Some time ago the Detroit Copper Mining Co. had to decide the question whether it would pay to re-treat slime-tailings, and several machines were tested in order to ascertain the type of construction which would give the greatest saving. In previous tests on the ore of this company, several tables proved far inferior to the Frue vanner, and for this reason they did not receive any serious consideration. In the late tests, however, one table was found to surpass the vanner in the percentage of saving made. The Frue vanner used in these tests had been adjusted approximately to the slope which the slime-vanners in the concentrator of the Detroit Copper Co. had at that time, about 2 in. between the inside of the wooden posts, or 0.272 in. per ft. It had not been determined, however, if this slope was the most economical one for treating the slime-feed in the mill, or for re-treating the tailings from the slime-vanners, and I was instructed to ascertain by test-runs, under different conditions, the best adjustments for the work to be done. In order to determine if a change in the adjustments of the machines would improve the results, it seemed sufficient to continue running the tests between the vanner and the table before mentioned, keeping the feed and all other conditions the same, but varying the adjustments of the vanner from run to run. I employed a vanner-man of long experience to operate these machines, who expressed the opinion that a slope between 28 and 2.5 in. between the posts (0.323 and 0.340 in. per ft.) would give the best results. Some runs made this way gave a considerable improvement. While for a slope of 2 in. (0.272 in. per ft.) the saving had been only 40.5 per cent. of that of the table, it was 49.2 per cent. for 2.5 in. (0.340 in. per ft.) slope. A decrease of

Jan 1, 1910

By Charles R. Austiuon, Carll H. Samans

It has been known for several years1 that certain alloys of the Konal type, containing commercial cobalt (99.32 per cent Co and 0.42 per cent Ni) and varying amounts of ferrotitanium, exhibit very high tensile strengths at 600°C. However, their proportional limits and yield pints are relatively low. published data' for the alloys quenched in water from 950°C. and then tested at 600°C. are summarized in the upper section of Fig. I. Similar specimens that were aged at 650°C. for 72 hr., after quench and prior to testing at 600°C., gave the result; shown in the: lower section of Fig. I. In the present work the investigation of the properties of these materials has been extended by a study of the tensile deformation characteristics in moderately longtime tests at 600°, 700' and 800°C. after quenching from various solution temperatures. In addition, some metallographic studies have been made to correlate with the deformation tests, and to obtain information on the quasi-binary system cobalt-Fe2Ti. The material, very kindly furnished by the Westinghouse Electric and Manufacturing Co. in the form of hot-swaged rods of 3/16-in. diameter, analyzed as shown in Table 1. The alloys, when originally made up,' were based upon an iron-titanium compound, Fe3Ti, as indicated by the work of

Jan 1, 1941

By Carll H. Samans, Charles R. Austiuon

It has been known for several years1 that certain alloys of the Konal type, containing commercial cobalt (99.32 per cent Co and 0.42 per cent Ni) and varying amounts of ferrotitanium, exhibit very high tensile strengths at 600°C. However, their proportional limits and yield pints are relatively low. published data' for the alloys quenched in water from 950°C. and then tested at 600°C. are summarized in the upper section of Fig. I. Similar specimens that were aged at 650°C. for 72 hr., after quench and prior to testing at 600°C., gave the result; shown in the: lower section of Fig. I. In the present work the investigation of the properties of these materials has been extended by a study of the tensile deformation characteristics in moderately longtime tests at 600°, 700' and 800°C. after quenching from various solution temperatures. In addition, some metallographic studies have been made to correlate with the deformation tests, and to obtain information on the quasi-binary system cobalt-Fe2Ti. The material, very kindly furnished by the Westinghouse Electric and Manufacturing Co. in the form of hot-swaged rods of 3/16-in. diameter, analyzed as shown in Table 1. The alloys, when originally made up,' were based upon an iron-titanium compound, Fe3Ti, as indicated by the work of

Jan 1, 1941