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By J. C. Wright

A practical program of education for workers of the mining industries is being formulated by the Federal Board for Vocational Education in cooperation with the states in which this industry is a dominant one. The Federal Government last year passed a law, known as the Federal Vocational Education Act, that provides for the promotion of vocational education and for cooperation with the states in the promotion of such education in agriculture and trades and industries and in the preparation of teachers of vocational subjects. More than 1,000,000 men in the continental United States are employed in mines and quarries. The following table, from the 1910 Census Report, shows the relative importance of mining and manufacturing pursuits: Manfactures Mines and quaRRies Employees.......... 7,405,313 1,109,410 Capital............. $18,428,269,706 53,380,525,841 Value of product..... $20,672,051,870 $1,238,410,322 Of the more than 1,000,000 men employed in mining, almost three-fourths are engaged in the production of bituminous and anthracite coal; one-fifth, in the production of copper, iron, lead, zinc, and other metals; and the remainder in getting out structural and miscellaneous materials from quarries. The meeting of the needs of the coal-mining industry, therefore, is the problem of greatest importance in the point of numbers to be served. Moreover, the dangers in coal mining are greater than in metal mining. Practically all states have set up educational and technical standards for the position of foreman, boss, superintendent, etc., and every applicant for these positions must have a certificate of competency. The qualifications, too, are becoming increasingly difficult in the field of technical information. In some states, every miner must have a certificate, which is issued by the mine examining board only after an oral examination. Did these state laws not call for definite and comprehensive plans for the vocational education of men in the mining regions, the conditions

Jan 1, 1919

By R.K. Matheson

The Nivloc mine is 9 miles west of Silver Peak, Esmeralda County, Nevada. It has been operated by Desert Silver, Inc., since the summer of 1937. The cyanide mill treats 190 tons of silver-gold ore per day. In March of this year the main operations were about 2000 ft. west of the main shaft, which served as the downcast air passage for the mine. The mechanical blowers in use were sufficient to ventilate the west end at that time, but it was foreseen that as the mining area was extended ventilation would be an increasingly difficult problem, especially during the summer months. It was decided, therefore, to continue a raise from a point 100 ft. above the highest mining level (440-ft. level) to the surface; a distance of 443 ft. at approximately 65°. From this point 40 ft. of drifting would be necessary to connect with a shallow shaft on the surface. (Fig. I.) All the raise was serviced from the 440-ft. level; costs and progress obtained in the initial 100 ft. are included in the following pages. Choice of Raise A raise in this locality would have four objects: (I) to provide an ample supply of fresh air to the working places; (2) to explore the vein from the highest mining level to the surface; (3) to provide waste fill for stopes in the area, and possibly provide a transfer for waste from the surface; (4) to provide an emergency exit in the west end of the mine. Before the mining of an ore shoot in the eastern part of the mine, a similar raise had been driven for purposes of ventilation and exploration. This raise and subsequent mining showed the vein to be noncommercial for a distance of about 250 ft. below the surface. However, in the area under operation, old surface workings showed occasional values. Thus from a development standpoint the raise was an excellent prospect. The raise chosen was a three-compartment, stulled raise having two chutes and a manway. The manway inside timber measured 4 ft. 8 in. by 4 ft. 6 in.; it carried air and water pipes, a ladderway and a slide for a timber skip. The two chutes would be used for drawing off the muck and the cleaning of the bulkhead, as well as for ore and waste passes, which would be especially useful for lateral exploration from the raise. The fairly large cross section of the raise (6 by 15 ft.) would provide a large quantity of waste for backfilling, and more could be obtained from a glory hole on the surface. Time Schedule As speed in completing the raise was the main requisite, three shifts were considered, but two shifts were decided upon when it was found that three would necessitate considerably larger fans than those available for clearing the raise. The main diffi-

Jan 1, 1943

By R. K. Matheson

The Nivloc mine is 9 miles west of Silver Peak, Esmeralda County, Nevada. It has been operated by Desert Silver, Inc., since the summer of 1937. The cyanide mill treats 190 tons of silver-gold ore per day. In March of this year the main operations were about 2000 ft. west of the main shaft, which served as the downcast air passage for the mine. The mechanical blowers in use were sufficient to ventilate the west end at that time, but it was foreseen that as the mining area was extended ventilation would be an increasingly difficult problem, especially during the summer months. It was decided, therefore, to continue a raise from a point 100 ft. above the highest mining level (440-ft. level) to the surface; a distance of 443 ft. at approximately 65°. From this point 40 ft. of drifting would be necessary to connect with a shallow shaft on the surface. (Fig. I.) All the raise was serviced from the 440-ft. level; costs and progress obtained in the initial 100 ft. are included in the following pages. Choice of Raise A raise in this locality would have four objects: (I) to provide an ample supply of fresh air to the working places; (2) to explore the vein from the highest mining level to the surface; (3) to provide waste fill for stopes in the area, and possibly provide a transfer for waste from the surface; (4) to provide an emergency exit in the west end of the mine. Before the mining of an ore shoot in the eastern part of the mine, a similar raise had been driven for purposes of ventilation and exploration. This raise and subsequent mining showed the vein to be noncommercial for a distance of about 250 ft. below the surface. However, in the area under operation, old surface workings showed occasional values. Thus from a development standpoint the raise was an excellent prospect. The raise chosen was a three-compartment, stulled raise having two chutes and a manway. The manway inside timber measured 4 ft. 8 in. by 4 ft. 6 in.; it carried air and water pipes, a ladderway and a slide for a timber skip. The two chutes would be used for drawing off the muck and the cleaning of the bulkhead, as well as for ore and waste passes, which would be especially useful for lateral exploration from the raise. The fairly large cross section of the raise (6 by 15 ft.) would provide a large quantity of waste for backfilling, and more could be obtained from a glory hole on the surface. Time Schedule As speed in completing the raise was the main requisite, three shifts were considered, but two shifts were decided upon when it was found that three would necessitate considerably larger fans than those available for clearing the raise. The main diffi-

Jan 1, 1943

By Donald J. Drinkwater, M. C. Fuerstenau

Many important contributions to the more fundamental aspects of mineral processing have been made this past year. Mular1 researched the flotation characteristics of pure zinc oxide and also samples that had been doped with Al+++ and Cu++. Correlation of the response of pure zinc oxide to oleate flotation with a calculated zpc (from solution equilibrium) suggests that adsorption of oleate ion is coulombic in nature. Doping the sample with Al+++ (which will make the ZnO more n-type) had a marked effect on the maximum value of pH at which 50% recovery was obtained at constant collector addition. That is, when 0.5 mole percent Al+++ was added, the pH for 50% recovery was reduced by about four units. Treatment with Cu++ was found to have little effect on the flotation response of zinc oxide, even for additions of one mole percent. Since the zero points of charge of the Al+++ doped powders were not measured, it was not possible to correlate flotation data of these samples directly with charge phenomena.

Jan 2, 1966

By N. E. Grosvenor

Several methods have been proposed for determining the tensile strength of rock. These have been tried out over the past several years, but the results have been erratic and of doubtful value. The tensile strength of rock is its weakest property, and this property apparently enters into all failure phenomena, whether underground or on the slope of an open pit. Therefore, it is essential that detailed quantitative information is known about the tensile strength of rock if the design engineer is going to select the proper pit slope, pit orientation, or orientation of underground workings so as to give the most economical and safest mining condition. This quantitative information may be used to orient the pit in such a manner as to allow steep slopes along two sides. This information possibly may be used to orient the direction of either underground workings or open pits so as to decrease powder consumption and to increase blasting efficiency. Hence it is of prime importance that a method of determining accurately the tensile strength of rocks be developed. The U. S. Bureau of Mines states' that no standard methods have been proposed by the ASTM for testing the tensile strength of rock. The method described in the Bureau publication is said by them to be unsatisfactory. A similar test that has been used in our laboratory in place of the USBM test is shown in Figs. 1 and 2. h this test the core is placed in a set of grips that has been designed for one particular size core. The grips are designed to distribute the stress around the core if they are tightened correctly. A universal joint or a flexible cable may be used in conjunction with the apparatus to prevent bending or twisting of the specimen. The main difficulty encountered with the system shown is the failure of the specimen within the grips. The possible cause of the failure within the grips may be due to a stress concentration set up by the uneven tightening of the grips. This method can be used only for tests on cores, and even then has been found to be totally unsatisfactory because of the failure within the grips. The U. S. Bureau of Reclamation has been using a test similar to the sketch shown in Fig. 3. Sulfur is cast around each end of the specimen to form a bearing surface and then a ring grip is used to apply a tensile stress. The grips are in contact with the sulfur cast and do not touch the specimen. This method is unsatisfactory, apparently because a large number of the specimens appear to break within the sulfur cast. Professor Wuerker of the University of Illinois uses the test shown in Fig. 4.' Professor Wuerker states2 that this method has given definite and reproducible results with the weakest as well as with the strongest rocks. The test is a supplement in all cases where cores cannot be obtained. The principle and equipment are the same as for the test for tensile strength of hydraulic-cement mortar. Soft

Jan 1, 1961

By Geo. Otis Smith

Twelve years ago, the Director of the United States Geological Survey addressed to the Secretary of the Interior a letter calling attention to the government's need for liquid fuel for naval use and pointing out that the rate of increase in demand was more rapid than the increase in production.' This letter, in a way, inaugurated the policy of public oil-land withdrawals, which was well founded in its primary purpose of protecting the oil industry and highly desirable in its immediate effect of checking the over-development of that day in California. Unfortunately, however, through delays in legislation, this policy may be regarded now as having outlived both its intent and its usefulness. In 1908, the country's production of oil was 178,500,000 bbl., and there was a surplus above consumption of more than 20,000,000 bbl. available to go into storage. In 1918, 10 years later, the oil wells of the United States yielded 356,000,000 bb1.—nearly twice the yield of 1908—but to meet the demands of the increased consumption more than 24,000,000 bbl. had to be drawn from storage. Nor is this all of the brief comparison. In 1918, our excess of imports over exports of crude petroleum was nearly 33,000,000 bbl. whereas in 1908 we exported 3,500,000 bbl., which was net, as we had not begun to import Mexican oil. In this period, the annual fuel-oil consumption of the railroads alone has increased from 16,871,000 to 36,714,000 bbl.; the annual gasoline production from 540,000,000 gal. to 3,500,000,000 gal. This record may be taken not only as justifying the earlier appeal for Federal action, but as warranting deliberate attention to the oil problem of today. Need of Future Supply The position of the United States in regard to oil can best be characterized as precarious. Using more than one-third of a billion barrels a year, we are drawing not only from the underground pools but also from storage, and both of these supplies are limited. In 1918, the contribu-

Jan 1, 1921

By Carl Henrich

While the size of the blast-furnaces used for smelting leadand copper-ores has constantly increased, during late years, the manner of removing the slag from the furnace to the slag-dump has (until quite recently, at least) practically remained the same. The only appliance in practical use has been the two-wheeled slag-pot, holding about 300 to 400 pounds of slag, or as much as one man could manage to pull or push on a smooth, hard surface. When these slag-pots on wheels were first introduced, the daily capacity of a lead or copper blast-furnace hardly reached 30 tons of smelting-charge. With furnaces of this size, and even when the capacity had been doubled, this kind of slag-pot answered well enough. All the improvements made were in the construction of these pots and not in an increase of their size, which was limited by the weight which one man could handle. But the size of the furnace went on increasing. The daily consumption of one furnace rose gradually to 75 tons, 100 tons, and even more. The writer had occasion to construct some of these furnaces, which, in their day, were among the largest, if not the largest, of the water-jacketed kind then in existence. A water-jacketed furnace will always smelt more in the same time than a furnace with brick walls. Already, in those days, the slag-removal from these enlarged furnaces received due attention by the writer. The removal of 60 tons of slag in 24 hours, in pots holding only 300 pounds each, will require 400 pots to be filled in 24 hours; and when the amount.of slag increases to 90 tons, 600 pots will be required. That is to say, in the last case, 25 pots in each hour, or 1 pot each 24 minutes, will have to be put under the slagspout, filled, pulled away, and replaced with another. This will involve too much hurry for any degree of comfort, convenience

Jan 1, 1896

By Charles H. White

Methods in colorimetry are based on the assumption that the intensity of the' color of a definite volume .of solution is directly proportional to the quantity of the color-producing substance present. In the preparation and examination of colored solutions there are three ways in which the depth of color may be varied. These variables are : (1) the quantity of' the coloring matter; (2) the quantity of the solvent, or the dilution ; and (3) the thickness or depth of that portion of the solution examined. With these three variables as a basis, three general methods in colorimetry have been devised, all of which have been used in the determination of carbon in steel. In the first method, the quantity of the solution is kept the same for both the unknown steel and the standard, and the sections compared are of equal thickness, while the quantity of the standard is varied until the color obtained is of the same intensity as that produced by a definite quantity of the unknown steel. In laboratories where this method is used there is prepared and kept for permanent use a series of solutions with varying amounts of the standard, representing the percentages of carbon from the lowest to the highest demanded in that laboratory. These standard solutions are kept in bottles or test-tubes of equal diameter, and are arranged in series from the faintest to the most intense color, so that it is a simple matter to find the place of the unknown steel in the series, when the solution is of a definite volume and is contained in a tube of the same diameter as the standard tubes. In the second method, equal quantities of the standard and of the unknown steel are dissolved, the standard is diluted to a definite volume, and the solution of the unknown steel is di-

Jan 1, 1907

New Jersey Zinc Co. - The Mines Franklin-Sterling Friedensville Development Austinville Ivanhoe Development Eagle Mine Hanover

Jan 12, 1953

THE opening session of the Institute was held at the rooms of the American Society of Civil Engineers, No. 4 East 23d Street, Vice-President R. W. Raymond in the chair. After a short introductory address by the chairman, the following names were presented by the Council for election as members and associates of the Institute. They were unanimously elected : MEMBERS. Abbott, James W., - Lake City, Col. Abbott, J. J., Jr., - Lake City, Col. Bartlett, J. C., - Cambridge, Mass. Billings, G. H., - Boston, Mass. Blake, Frank C., - Lafayette College, Easton, Pa. Dewey, Fred. P., - Lafayette College, Easton, Pa. Douglass, Samuel T., - Ann Arbor, Mich. Emmons, S. F., - New York City. Engle, George U., - Philadelphia. Harrington, Dr. Bernard J., - Montreal, Canada. Hayden, Wm. B., - Columbus, O. Hill, Leslie C., - Capelton, Quebec, Canada. Land, W. J., - Atlanta, Ga. May, Wm. A., - Scranton, Pa. Mills, James E., - St. Louis, Mo. Moore, Charles W., - Montezuma, Ind. Page, William Byrd, - Lexington, Ky. Putnam, Douglas, Jr., - Ashland, Ky. Richards, Henry, - Dover, N. J. Roberts, Percival, Jr., - Philadelphia. Sackett, S. G., - Brooklyn, N. Y. Schley, Wm. S., - Atlanta, Ga. Setz, Gustav, - Mine La Motte, Mo. Talbutt, John H., - Lexington, Ky. Torrey, Herbert G., - U. S. Assay Office, New York City. Wernlitz, Fred. W., - Cartersville, Bartow Co., Ga. Wiestling, Geo. B., - Mont Alto, Franklin Co., Pa.

Jan 1, 1877

By H. Scott

Since Bohler discovered, in 1903, on cooling certain alloy steels, the phenomenon of a new and lower temperature transformation than the usual Ar 3-2-1 obtained by increasing the maximum temperature to which the material was heated, much has been published1 connecting this phenomenon with a large number of dissimilar steels of high alloy content. From the fact that the transformation divides itself, taking place at two widely separated temperatures, it has been called a "split transformation." The significant facts established by recent investigators2 are: that when the transformation occurs at the higher temperature Ar1, troostite or a decomposition product is formed, and that when the transformation occurs at the lower temperature Ar", the resulting structure is martensite. The terminology Ar' and Ar" adopted here is that of Portevin. Previous Investigations Reviewing the work on this subject published in a recent issue of Revue de Metallurgie and referring in particular to his statement that martensite is a solution of carbide in alpha iron, Le Chatelier says :4 "How then can a theory already 20 years old demand new investigations? The reason for it is that we have not succeeded in proving directly the real presence of the transformation of iron during the very short duration of the quenching. The fall of temperature takes place at the rate of several hundred degrees per second and the observation of

Jan 1, 1920

By D. J. Brown

Froth flotation is now used to recover millions of tons of fine coal,(-1/2 mm), every year. For example, in Great Britain, 4-million tons of flotation concentrate was produced in 1960, about 3% of the mechanically cleaned out- put and 2% of the total salable output. Coal flotation is not so widely practiced in the U. S. as it is in Europe since much American coal occurs in thick seams of high grade and does not require cleaning. However, in the U. S., as elsewhere, the increasing mechanization of coal mining machinery has lead to a lowering in size and grade of product from the mine and to an increasing need for cleaning of coal to improve the grade before sale. With an increasing proportion of fines in the product, froth flotation has assumed a greater importance as a cleaning process for coal. In the last three years, over 40 flotation plants have been put into operation in the Appalachian area of the U. S.80 Fine coal recovered by means of froth flotation may be used as a powdered fuel alone or mixed with coarser grades. The removal of this fine coal from washery water makes possible the recirculation and re-use of the water and also leads to improved and more consistent operation of jigs and other separating machines which function less effectively when there is a build-up of fine solids in the washery water. Finally, froth flotation is of assistance in the production of a clarified effluent which may be discharged from the washery without polluting local rivers, streams, or wells. The 1961 cost of installing a new coal flotation plant in Great Britain would be about $19,000 per ton of dry solids input per hour, including the cost of new buildings and coal and shale filtration and water clarification equipment. The figure for the gravity concentration of the large coal, the main plant, would be about $4,800 per tph. The operating cost for the new coal flotation plant would be about $1.10 per ton. Even at this level of cost the cleaning operation is economic in Great Britain for coking coals. In the U. S., a froth flotation cost of 14$per ton of cleanfilter cake was quoted early in 1958 for one operation,l though admittedly this was a special case. Assuming a cake moisture con- tent of 30%, it is estimated that this cost is equivalent to a figure of 256 per ton of dry solids input, for the incorporation of coal flotation in an existing washery. A cost oh 416 per ton of concentrate was quoted76 in a recent analysis of flotation economics, this cost being brought up to $1.11 per ton after payment of royalty, welfare and association dues and taxes. Capital cost was $120,000 for a plant producing 133 tpd of concentrate. Before considering the treatment of coal by flotation it is appropriate to survey some of the properties of coal which are of importance in the process.

Jan 1, 1962

By Anson Hayes

THE quality of sheet and strip products made of rimming steel is closely related to the structure and chemistry of the ingots. The varia-tion in composition throughout the ingot, as affected by segregation, is of great importance, as are also the form and distribution of the voids that result from the gases evolved. It is because of this importance that the considerable literature in this field has accumulated. No attempt is made in the present paper to compile a bibliography of the excellent work that has already been done, as this has been adequately covered in the many excellent reports on the heterogeneity of steel ingots prepared by a joint committee and published by the British Iron and Steel Institute.1 From time to time, in any field of investigation new tools and new technique are developed, which make further progress possible. In the field with which this paper is concerned, this is true. The development of the vacuum-fusion method as a means of following oxygen segregation, and a technique for obtaining the volume and composition of the gases evolved during solidification, have opened the way to obtaining data by means of which some considerable progress has been realized in developing more of the detailed mechanisms of solidification and segregation in rimming-steel ingots.

Jan 1, 1938

By David Swan

No mystery shrouds the discovery of chromium. The indefatigable French chemist Vauquelin, in reporting his discovery of this chemical element to 1'Institut National in 1798, concluded his discourse with the statement: "I place before the eyes of the Institute a small quantity of chromium and several of its most re- markable combinations." Working with the so-called "red lead of Siberia, Vauquelin not only prepared numerous chemical compounds of the new metal, but he reduced the oxide to the metallic state. These chemical compounds possessed a variety of remarkably distinctive colors; so, acceding to suggestion, Vauquelin named the new element chrome (chromium), a derivative of the Greek word chroma, meaning color. Siberian "red lead" (crocoite) was not plentiful; however, it was not long before Vauquelin's prediction was realized that the chemical compounds of chromium would become useful in the arts when sources of chromium became more readily available. Although chrome-iron ore (chromite) became obtainable in France, the United States and elsewhere during the early part of the nineteenth century, it is interesting to note that from 1828 to about 1860, Maryland, Pennsylvania, and others of the eastern United States furnished almost the entire chrome-ore requirements of the world. Several countries each in turn have since dominated for certain periods the supply of chromite. In recent years, the largest tonnages have come

Jan 1, 1953

BEFORE me is a colored mining-claim map of the productive portion of the Robinson Mining district-sometimes called the Ruth-Kimberly area-that lies a few miles west of Ely, Nevada. It covers an area of 20 sq mi and at the end of 1954 had produced just under 2,000,000 tons of copper and about 2,000,000 oz of gold as a by-product. The three principal colors on the map and the meaning of each are: Red -Property of the Nevada Mines Division of Kennecott Copper Corporation, formerly known as the Nevada Consolidated Copper Company. Yellow -Property of the Consolidated Coppermines Corporation. Green --Consolidated Coppermines property under lease to Nevada Mines. The striking feature of the map is the crazy-quilt intermingling of the colors and the tendency of property lines to intersect ore bodies. For example, the westerly one-fourth of Nevada Mines' Copper Flat pit, source of most of the ore,up to date, is in Coppermines property. Nearly half of Nevada Mines' Veteran pit is on leased claims; and Nevada's new Deep Ruth shaft is on a right of way on ground originally owned by Coppermines. On the other hand, the original Ruth underground mine and the new Kimbley pit are on Nevada Mines property; and the Morris and Brooks opencut mines, the most important of Cop-

Jan 1, 1957

By Frederick T. Fischer

The New Jersey Zinc Company began a program of exploration in Middle Tennessee in 1964. The target horizon of the exploration project has been the Knox Dolomite which is a low-yield aquifer over nearly the entire Central Basin of Tennessee (1). With the exception of a few hundred acres in the Wells Creek and Flynn Creek disturbances (astroblemes?), the Knox Dolomite in Middle Tennessee is known only in subsurface. In January of 1969, New Jersey Zinc announced the discovery of a major zinc deposit near Elmwood, southeast of Carthage. This is on the northeastern flank of the Nashville Dome where the nearly flat-lying beds have an average eastward slope of about 50 feet per mile. The anticipated development of a mining situation has emphasized the significance of hydrologic relationships in the region as a whole and especially in the Elmwood area. These relationships affect several areas of concern which ma be generally grouped into two broad classes; i. e. , (1) Problems of a Public Nature - involving the prevention of the contamination of potable aquifers, the prevention of the contamination of surface waters, the prevention of the capture of presently utilized water, etc.; and (2) Problems of Mining Economics - involving pumping capacity and costs, the possible need for equipment of special composition, underground environmental conditions, etc. Exploration drilling in Middle Tennessee has contributed additional information about the Knox aquifer and

Jan 1, 1970

By Jacob T. Wainwright

AS a means for increasing the efficiency in modern blast-furnaces by supplying to them blast of a much higher temperature than is now possible, the writer offers as a suggestion a modification in the construction of regenerative hot-blast stoves. Stated briefly, the plan proposed is to pre-heat both the gas and the air for its combustion in a manner similar to that generally adopted in Siemens openhearth furnaces. After thus carrying the temperature of the stove to a point much higher than that common in hot-hlast stoves, the cold blast passes through it. Without entering into the details relating to the reversing valves and the system of flues, the general plan may be outlined as follows, referring to the accompanying drawings: The valves are so placed in heating up the stove that the blast-furnace gas enters at B, Fig. 1, and the cold air for combustion at A. Ascending in flues, F and E, respectively, combustible gas and air meet and mix in the combustion-chamber, I. Flowing through K, they descend through the flues, G and H, escaping through the chimney. In this manner the walls of the combustion. chambers, I and L, are highly heated, and the right-hand side of the stove is carried to a higher temperature. By reversing the valves, blast-furnace gas enters at C and cold air for combustion at D. Both are pre-heated in their ascent through the flues, G and H. They burn in the combustion-chamber, L, and the hot gases of combustion, flowing through K, I, E and F to A and B, heat the left-hand side of the stove. In this manner the stove is carried to a temperature greater than that of those now generally in use, in which air and gas are burned without preliminary pre-heating. In the plan proposed, that side of the stove will, of course, be relatively cooler through which cold air and furnace-gas last passed in its ascent to its combustion-chamber. After having sufficiently heated the stove in the manner outlined, the cold blast is admitted through all the flues, A, B, C and D. It ascends through the highly-heated regenerative material to the chambers at the top and descends through the flue, P, Fig. 2 (dotted

Jan 1, 1889

By H. J. Snyder

Both the room-temperature and 600°F (315°C) primary solid-solution phase boundaries for the silver comer of the Ag-Cd-In temary system have been determined using X-ray diffraction and metallo-graphic techniques. Solid solubility of indium in silver at room temperature and 600°F (315°C) is shown to be between 20.3 and 21.7 wt pct, and corrects the previously extrapolated value of 18 wt pct appearing in the literature. The solid solubility of cadmium in silver at room temperature, not previously established, is shown to be between 34.9 and 41.3 wt pct. The solid solubility of cadmium in silver at 600°F (315°C) is confirmed as 43.2 wt pct. The shape of the primary phase boundary line for the subject temary system appears straight at room temperature and curves con-vexly towards the solvent (silver) comer at 600°F (315°C). THIS investigation was undertaken to initiate development of the Ag-Cd-In ternary diagram by establishing the primary solid-solution phase boundary in the silver corner of the diagram. At the present time, a diagram for this ternary system is not available, while this system appears to offer potential for a variety of uses. One alloy in the system has already proven to be useful as a control rod material in nuclear reactors.1"3 To encourage additional exploration of the potential of the system an understanding of its ternary diagram would be beneficial. It is towards this end that the present study is directed. LITERATURE SURVEY In undertaking this study, a comprehensive literature survey was first conducted to find information that would be useful for guidance. Useful knowledge did result from investigating the respective solubilities of indium and cadmium in silver as indicated in the binary constitutional diagrams of the systems Ag-In and Ag-Cd and also reviewing possible shapes of ternary primary phase boundaries. Weibke and Eggers~ used metallographic techniques to establish the system In-Ag as illustrated in Fig. Is This diagram shows that silver in the solid state is capable of dissolving approximately 18 wt pct In. This solid solubility of indium in silver holds true from room temperature up to 1276°F (691°C), then decreases continuously up to 1'761°F (960.5"C), the melting point of the silver. The solid-solubility limit of indium in silver established by Weibke and Eggers has been confirmed by Hume-Rothery and Reyno1ds. G Durrant' used a method which consisted of microscopic examination of quenched specimens to establish the Cd-Ag system as displayed in Fig. 2.' The primary phase boundary of Fig. 2 proved to be in good agreement (ie., within 1 wt pct) with other studies*'1' of the Ag-Cd system made using X-rays and metallog-

Jan 1, 1968

By James Terry Duce

Production of oil in Bahrein for the war years was as follows: YEAR BBL PER YEAR Prior to 1939 . . . . . . . . . . . . . . 22,287,055 1939 . . . . . . . . . . . . . . . . . . . . . . 7,588,554 1940 . . . . . . . . . . . . . . . . . . . . . . 7,073,919 1941 . . . . . . . . . . . . . . . . . . . . . . 6,794,157 1942 . . . . . . . . . . . . . . . . . . . . . . 6,240,627 1943 . . . . . . . . . . . . . . . . . . . . . . 6,571,609 1944 . . . . . . . . . . . . . . . . . . . . . . 6,713,795 1945 . . . . . . . . . . . . . . . . . . . . . . 7,308,938 Total to Jan. 1, 1946 . . . . . . . . 70,578,654

Jan 1, 1946

Duval Corporation, the copper -producing subsidiary of Pennzoil United, Incorporated, operates two plants south of Tucson, Arizona: Esperanza and Sierrita. The newest, the Sierrita mine and mill, came into full production in 1971 at a rate of 72,000 tpd (now 87,000 tpd) which at that time made it the largest designed copper plant in the world. In 1968, the largest milling units then available were selected; these included fourteen 16$ ft by 19 ft long ball mills (5 m by 5. 9 m), each equipped with a 3,000 hp motor. The primary crushing plant contains two 60 in. by 89 in. Allis- Chalmers gyratories; each is constructed for two-side dumping by 120- and 150-ton trucks. Access to the maintenance areas to service the drive motors, eccentrics, apron feeders, and speedup conveyor is through hatches between the crushers. This area also contains a good- sized elevator and a stairway. The rather shallow dump pocket, the 300 -ton crushed ore pocket, and the overall design make for a fine, compact arrangement for two 60 in. crushers installed in a total height of 89 ft (27. 13 m) from dump level to top-of-concrete at the lowest level beneath the collecting or speedup belt. This efficient layout is reflected in the short changeout time for mantle and concaves (8 and 30 hours respectively) and the low crushing cost of $0.0416 per ton (first half of 1975). The crushers are installed on 64 ft (19.5 m) centers (Bougainville and Palabora, both with 54 in. crushers, are 55 and 43 ft or 16.76 and 13. 1 m respectively). The switchgear room is located between the truck dumping positions on one side of the installation and the hydroset units on the other side, with dust collecting equipment mounted over the latter. Protection of these units from rocks falling from the trucks is provided by concrete walls. Truck dumping is controlled from a central 25 ft high (7.6 m) control room, which permits fine observation of operations. Removal of oversize boulders from the dump pockets is done with a hydraulic grapple or a rock hook or sling attached to the auxiliary crane hook. Spare mantle stands and a maintenance room with a zinc melting furnace are located at one end of the surface facilities. The maintenance room is equipped with an access door for the crane, which is equipped with 80- and 20-ton hoists. The two 84 in. (2. 135 m) apron feeders receiving the 6 in. (152 mm) crushed rock discharge onto a 72 in. (1. 83 m) collecting and wear belt operated at a speed compatible with the 60 in. (1.524 m) takeaway belt running at a speed of 630 fpm (192 mpm). This belt feeds a 54 in.

Jan 1, 1978