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By Donald E. Garrett

The processing of brines from Searles Lake, California, yields one of the world's largest supplies of borax and boron chemicals, second probably only to the massive deposit at Boron, California. Two companies produce from this deposit, the American Potash & Chemical Corp. and the West End Chemical Co., a division of the Stauffer Chemical Co. Two processes are employed in recovering borax from the brines; one is used by both companies and consists of carbonating the brine to form sodium bicarbonate, removing this, and then cooling the liquor to crystallize a crop of borax. Only American Potash uses the second process, evaporating the brine to concentrate it and remove (crystallize) much of its NaCl, Na2CO3, Na2SO4, and NaLi2PO4 content, quick cooling the remaining liquor to remove KCI, and then seeding the filtrate to remove borax. Both methods will be briefly described. Searles Lake is a large evaporite body formed during the most recent two Sierra Nevada ice stages, the Tahoe and the Tioga. Runoff waters from the Sierra Nevada Mountains collected in the large Owens Valley drainage basin and from there overflowed into the Indian Wells basin and then into Searles Lake. In each Lake, evaporation took place. To a large extent the Owens and Indian Wells lakes acted as settling basins, for much of the mud, and insoluble calcium and magnesium precipitates were deposited there. Recent volcanic activity in the mountains to the north and east of Owens Lake had scattered considerable mineral-rich volcanic ash and debris throughout the drainage area, and several hot springs were formed that contained boron and other salts. The present flow rate and concentration of several springs, and the known age of at least the Tioga zone, can fairly well account for all of the boron present in Owens and Searles Lakes.1 The tungsten content can be accounted for in a similar manner.2 Most of the salts thus originally present in the fresh water runoff-from ice, snow, and rain-accompanying these two ice stages were ultimately deposited in Searles Lake. During the Tahoe period, from 25 to 40 ft of massive, mixed salts were deposited, along with six interspersed layers of mud (four inches to two feet thick) representing comparatively brief interglacial periods. Before the final desiccation of the lake, from 10 to 14 ft of new mud was deposited, representing the Tioga stage. The crystals subsequently deposited were from 70 to 90 ft thick, and form the present lake surface. The extent of the salt body is about 34 square miles, of which 12 square miles are exposed salt; the remaining area is covered by recent alluvial deposits (chiefly mud). The present lake is generally dry on the surface but floods to a shallow depth following heavy rains. The crystal body of both zones consists of continuous clusters of salts, and the voids between crystals (about 45 pct of the total volume) are occupied by brine. The level of this brine never recedes more than several inches below the salt surface. The brine pumped to the plants for processing is generally saturated with all of the solid phases present in that area, and thus is a complex mixture containing about 35 pct solids (see Table 3). The brine composition varies somewhat at different areas of the lake and from upper to lower structures, so wells are drilled in the areas and to the depths desired. A casing is cemented in the hole to within a short distance from the bottom, and pumps with about 14-ft suctions are mounted on the casing to pump the brine into the plants. The evaporation process of the American Potash & Chemical Corp. treats upper structure brine, evaporating all of the water from

Jan 1, 1960

By A. H. Larson, L. G. Twidwell

The influence which Au, Ag, Sb, Bi, Sn, and Cu have, both individually and collectively, on the activity coefficient of sulfur in liquid lead at 600"C zuas studied by circulating a H2S-Hz gas wlixture over a specific lead alloy until equilibrium was attained. Subsequently, the H2S concentration in the equilibrium gas mixture and sulfur concentration in the condensed phase were deterruined. The elements gold, silver, and antinzony (above 8 at. pct) increased the activity coefficient of sulfur. Bismuth had no apparent effect. Tin (above 3 at. pct) and copper decreased the coefficient. The influence of an individual element, i, on sulfur is best reported as the interaction parameter, riS, which is defined as The values o these first-order interaction zus are: ESzu = —55.0. These interaction parameters are used to predict the activity coefficient of sulfur in six fouv-component alloys and one seven-component alloy. Comparisons are made with direct experimental determinations. INTERACTIONS in dilute solution have been studied by many investigators. Most of the experimental work has been confined to solute-solvent interactions in simple binary systems and solute-solute interactions in ternary systems. Dealy and pehlke"~ have summarized the available literature on activity coefficients at infinite dilution in nonferrous binary alloys and have calculated from published data the values for interaction parameters in dilute nonferrous alloys. Interaction parameters are a convenient means of summarizing the effect of one solute species on another in a given solvent. Only a few investigators have studied interactions of the nonmetallic element sulfur in a metallic solvent. They are as follows: Rosenqvist,~ sulfur in silver; Rosenqvist and Cox,4 sulfur in steel; chipman, sulfur in alloy steels; Alcock and Richardson,% ulfur in copper alloys; Cheng and Alcock,' sulfur in iron, cobalt, and nickel; Cheng and ~lcock,' sulfur in lead and tin. The only reported work on the Pb-S system in the dilute-solution region is that of Cheng and Alcock.' Their investigation involved a study of the solubility of sulfur in liquid lead over the temperature range 500" to 680°C. The results may be summarized by the following relationship: S (dissolved in lead) + Pb(1) = PbS(s) log at. %S = -3388/T + 3.511 Experimentally, it was found that Henry's law was valid up to the solubility limit of sulfur in lead, i.e., at 600°C up to 0.43 pct. Their investigation did not include the study of sulfur in lead alloys. More accurate calculations could be made in smelting and refining systems if activity coefficients of solute species could be accurately predicted in complex solutions. One of the objectives of this study was to compare the experimental data with the values calculated from the equations derived from models for dilute solutions proposed by wagner9 and Alcock and Richardson. A temperature of 600°C was chosen as the experimental temperature to attain reasonable reaction rates and to minimize volatilization of the condensed phase. EXPERIMENTAL Materials. The Pb, Au, Ag, Sb, Bi, Sn, and Cu used for preparation of the alloys were American Smelting and Refining Co. research-grade materials. All were 99.999+ pct purity except the antimony and tin which were 99.99+ pct. The initial alloys prepared for this study consisted of twenty-one binary alloys, eleven ternary alloys, and one six-component alloy. The constituent elements were mixed for each desired alloy and were placed in a crucible machined from spectrographically pure graphite. The crucible was placed in a vycor tube which was evacuated with a vacuum pump and gettered by titanium sponge at 800°C for 8 to 12 hr. After the gettering was completed, the chamber containing the titanium was sealed and removed. The remaining sample chamber was placed in a tube furnace at 800°C for 2 hr and quenched in cold water. The final operation consisted of homogenization of the alloy for 1 to 2 weeks at a temperature just below the solidus for the individual system. The resulting master alloys were sectioned into small pieces and a random choice made for individual equilibrations. Cobalt sulfide (Cogs8) used to control the gas atmosphere in the circulation system was prepared by passing dried HzS for 24 hr over a Co-S mixture heated to 700°C in a tube furnace. This material was then mixed with cobalt metal to give a two-phase mixture which, when heated in hydrogen to a particular temperature, produced a desired H2S/H2 gas atmosphere in the circulation system. A Cu2S-Cu mixture also used in this study was prepared in a comparable manner. Apparatus for Equilibrium Measurements. The experimental technique of this study required apparatus

Jan 1, 1967

By H. B. Pulsifer

One of the most common basic open-hearth furnace products is a simple carbon steel with a carbon range from 0.05 to 0.15 per cent. The material is widely used for sheets, tubes, bars, wire and the innumerable special objects of secondary fabrication. The properties of the material vary over wide ranges, depending on the more exact composition and details of manufacture or treatment that affect its structure. The present paper is a review of those properties and structures that have been found by studying this steel as used in the cap screw industry. At first glance this might be thought a narrow application of the steel, but, since the industry uses the material in the cold-worked, annealed and heat-treated conditions, the title of the paper may be excused for its comprehensiveness. Yet, in studying the wide ranges of properties and structures that preclude a more specific title, there can be no suggestion of completeness or finality. Even a summary of a few years' work is no more than a preliminary survey of a field that will be fertile in scientific data and industrial products for many years to come. An outline of the structures will be given first, then a list of the previous contributions that are at hand and finally some details of the work and related problems, especially concerning this low-carbon steel in the surface-hardened condition. Wire Used in Manufacture of Cap Screws The wire that is used for making cold-upset and roll-threaded cap screws has been cold-drawn from hot-rolled rods that were pickled and limed. The common sizes used in the industry range from 0.185 to 0.750 in. dia., the exact diameters depending on the plant practice and having the usual tolerances of 0.002 in. for the smaller sizes and 0.003 in. for the larger sizes. The carbon range, by pit analysis, is usually between 0.07 and 0.12 per cent. Manganese should be within the usual 0.30 to 0.50 per cent. range. Sulfur rarely exceeds 0.045 per cent. and phosphorus seldom exceeds 0.030 per cent. Silicon is usually below 0.03 per cent. but in the silicon-killed steels may be in the customary 0.10 to 0.25 per cent. range. Rimmed steel is not considered most suitable for cap screws that are cold-headed or for parts that are to be surface-hardened.

Jan 1, 1931

NEW MEMBERS The following list comprises the names of those persons who became members during the period Feb. 10 to Mar. 10, 1915: Members BANKS, HAROLD PURDY, Min. Engr 61 Broadway, New York, N. Y. BARAGWANATH, JOHN G., Min. Engr., Morococha Mining Co.; Morococha, Peru. BEESON, JOSEPH JOSIAH, Min. Geol Stanford University, Cal. CALLOWAY, ALFRED W Pres., Davis Coal & Coke Co., Cumberland, Md. DEETRICK, JAMES W., Mgr., Republic Iron & Steel Co., 826 Pennsylvania Ave., Youngstown Ohio.

Jan 4, 1915

By R. Steinitz, L. Seigle, R. Resnick

THE rates of growth of the carbide phases MC and M2C in the Ta-C and Cb-C systems have been measured at various temperatures and the results used to estimate the rate of carbon diffusion within the inter metallic phases. Sheets of high-purity tantalum (50 ppm Cb, <25 ppm Ti, 10 ppm Fe, 5 ppm Cr, <50 ppm Mo, W, and all other metals, 15-75 ppm O2, 10-30 ppm N2, and 10-30 ppm C) approximately 1/4 by 1/2 by 0.020 in. were heated to temperatures between 1800" and 2700°C in a carbon resistor furnace under an atmosphere of argon plus sufficient methane to maintain an excess of carbon at the sample surface. The thicknesses of the carbide layers, illustrated in Fig. 1, were measured with a filar micrometer and plotted logarithmically as a function of time. Fig. 2 is a typical plot for the TaC phase. The points may be represented within experimental error by lines with slope equal to 1/2; therefore the growth of the layers may be described by the usual expression X = kt1/2, where X is the layer thickness, t is the time, and k is a temperature-dependent constant. From these data and compositions at the phase boundaries, it is possible to estimate values for the diffusion coefficients of carbon in each phase, assuming that layer growth occurs by carbon diffusion alone, and that the dif-fusivity is not a function of composition within each

Jan 1, 1965

By F. C. Schora, H. P. Meissner

Iron ore from Cerro Bolivar, Segre&apos;, and Sierre Grande was reduced in fluid beds at about SOOT, using gas analyzing 20.5 pct CO, 41 pct Hz, and 38.5 pct N2. Except in the early stages of reduction, the rate of oxygen loss from the bed was directly proportional to the oxygen concentration in the bed, independent of gas composition as long as the gas was reducing, and little affected by temperature. The percentage reduction of the bed solids at any time was independent of particle size. I HE reduction of spheres or cubes of dense iron ore in a slowly moving stream of either pure H, or pure CO has been found to take place at distinct interfaces between well defined layers of iron and the iron oxides.&apos;y4 These interfaces generally penetrate the ore lump in topochemical fashion, remaining parallel to the original contours of the specimen. Thus in the earlier stages of reducing hematite, the specimen&apos;s core is Fe,O, surrounded by successive shells of the lower oxides, and finally by an exterior shell of iron. McKewan5 assumed that the rate limiting step occurred at the Fe-FeO interface, with reducing gas penetrating up to this interface through the external porous iron shell, and reduction of the higher oxides within the ore lump occurring by diffusion of iron through the dense FeO shell. He developed an analytical expression based on this model relating the amount of oxygen removed from

Jan 1, 1962

NEW MEMBERS The following list comprises the names of those persons who became members during the period of Apr. 10, 1918, to May 10, 1918. BALLARD, JOHN I Supt., Cable Mining Corpn., Cable, Mont. BURKMAN, JOHN M. F., Constr. Engr., U. S. Smelt. Refin. & Min. Co., Care -University Club, Mexico City, D. F. CALLEN, ALFRED COPELAND, Head, Mining Engineering Dept., West Virginia University Morgantown, W. Va. CHADWICK, LAWRENCE, Chem., Mt. Elliott Copper Co., Ltd., Selwyn, via Townsville, North Queensland, Australia. CODY, FRANK W., Skimmer, Reverberatory Furnace, Arizona Copper Co., Clifton, Ariz. COOLBAUGH, MELVILLE F., Prof. of Chem., Colorado School of Mines, Golden, Colo. DODGE, ALLAN WAYNE The Dodge Agency, 104 East Iron Ave., Salina, bans. DONNELLY, LEONARD G., Pet. Geol., Sinclair Oil & Gas Co., Sinclair Bldg., Tulsa, Okla.

Jan 6, 1918

NEW MEMBERS The following list comprises the names of those persons who became members during the period Mar. 1 to 15, 1914: Members ADAMS, JOHN H., Cons. Engr First National Bank Bldg., Birmingham, Ala. ALEXANDER, WILLIAM H., Miner 80 Maiden Lane, New York, N. Y. AUSTIN, EDGAR TAYLOR, Asst. Smelter Supt., Mond Nickel Co., Ltd., Coniston, Ont., Can. BAKER, DANIEL, JR., Asst. Treas., Standard Lime & Stone Co. 524 Equitable Bldg., Baltimore, Md. BALL, MAX WAITE, Geologist U. S. Geological Survey, Washington, D. C. BARREN, HARRY B., Supt. Blast Furnaces, Inland Steel Co., Indiana Harbor, Ind. BAUMANN, HENRY N., JR., Amalgamator 365 Lynn St., Seattle, Wash. BLACK,, THOMAS B., Min. Engr Care Vulture Mines Co., Wickenburg, Ariz. BRADT, HARLAN H 429 West 111th_ St, New York, N. Y.

Jan 4, 1914

By R. J. Marcotte, C. T. Eddy, R. J. Smith

The objectives of the investigation herein reported were to determine: (I) the S-curves for certain selected cast steels, (2) whether or not the published S-curves for wrought steels are satisfactory for use on cast steels of similar composition, and (3) to prepare hardenability curves of the selected cast steels. In order to make the final comparisons between cast and wrought steel more completely tangible, it was deemed advisable not only to determine S-curves for the steels submitted in the cast condition but also to roll some of the cast-steel test bars and determine comparison S-curves on such wrought material. This procedure eliminates the composition variable and makes it possible to reveal the differences resulting from plastic deformation of the cast structure in addition to providing the comparison with published data on wrought steels. Cast Steels The cast steels selected for the preparation of S-curves, or time-temperature transformation curves, were similar to S.A.E. series 1030, 2330, 4130 and 4330. Of these the 1030 was supplied by The Crucible Steel Casting Co., Cleveland, Ohio, and the other three by The Sawbrook Steel Castings Co., Lockland, Ohio. The steels were furnished in the as-cast condition. The manufacturers supplied the chemical analyses and the properties that could be obtained in the heat-treated condition—all of which is summarized in Table I. The metallographic and dilatometric procedures and equipment were in general standard. The specific technique employed was described in a report of the Steel

Jan 1, 1945

By Marvin L. Kay

In the spring of 1950, just two years after the inauguration of the domestic uranium program by the U. S. Atomic Energy Commission, the company entered the uranium mining and milling business through a newly formed subsidiary, Climax Uranium Co. This was the first venture into a new field of mining for Climax, which became the first old-time mining company, other than the vanadium producers, to become active in the embryo uranium industry. Twenty-five years earlier, Arthur Bunker, president of Climax Molybdenum, organized and operated the U. S. Vanadium Co. at Rifle, Colo. With this background, he viewed the budding industry of the Colorado Plateau with much interest. Upon its formation, the new subsidiary entered into a con- tract with the U. S. Government and the potential uranium-vanadium properties of Minerals Engineering Co. were assigned to Climax Uranium Co. Funds for the construction of a processing plant and working capital were advanced by Climax Molybdenum Co. The old sugar mill building of the Holly Sugar Co. in Grand Junction was leased to house the extractive plant. Construction of facilities, which were based upon a new process developed for the treatment of both uranium and vanadium ores, was completed and the mill was placed in operation early in 1951. Subsequent to that time several alterations were made and the capacity of the plant increased. In terms of recoveries, the plant has proved to be one of the most efficient on the Colorado Plateau

Aug 1, 1955

By W. B. Braden

This paper presents a suitable method for predicting gas-free oil viscosities at temperatures up to 500F knowing only the API gravity of the oil at 60F and the viscosity of the oil measured at any relatively low temperature. The API pravity and the one viscosity value are used as parameters to determine the slope of a straight line on the ASTM slanaord viscosity-temperature chart. Then, knowing the slope of the line and one point on the line, the vrscosities at higher temperatures can be determined. The line slope correlations were developed at I00 and 210F since viscosity data are frequently measured at these temperatures. A procedure is given for predicting line slopes from measurements at other tetnperatures. A nomogram is furnished for solving the relationship. The correlation has been evaluated at temperatures up to 5OOF for oils varyzng in gravity from 10 to 33 " API. The correiution is applicable only to Newtonian fluids. Comparison at 500F of true viscosities and those predicted from values at 100F shows an average deviation of 3.0 per cent (maximum deviation of 6.0 per cent). Predictions from the values at 21 0F for the same oils how an average deviation of 1.5 per cent (maximum deviation of 3.4 per cent). INTRODUCTION Correlations have been developed by Beal&apos; and by Chew and Connally&apos; for predicting viscosities of gas-saturated oils at reservoir conditions. Each of these correlations requires a knowledge of the solution gas-oil ratio and the viscosity of the gas-free oil at the reservoir temperature. For temperatures below 350F, measurements of the gas-free oil viscosities can be made easily using commercially available equipment. In thermal recovery processes, however, reservoir temperatures well in excess of 350F are encountered. Viscosity measurements at such conditions are more difficult and time consuming and require modification of existing equipment or the construction of new equipment. Measurements are further complicated by the difficulty of handling highly viscous oils associated with thermal recovery processes. Therefore, it is desirable to have a correlation which allows accurate prediction of viscosities at high temperatures. A commonly used technique for predicting viscosities at high temperatures is to measure the viscosities at two lower temperatures, plot the values on ASTM standard viscosity-temperature charts and extrapolate to the temperatures desired. If either of the values is slightly in error, the extrapolated value can be significantly in error. To justify an extrapolation, three points are actually necessary. This procedure can consume much time, particularly with heavy oils. Considering the cost of viscosity measurements, it would be desirable to eliminate the need for direct measurements by having correlations which would allow viscosity predictions from other physical or chemical properties. Beal1 investigated the possibility of correlating viscosity with oil gravity at temperatures from 100 to 220F. While showing that a general relationship exists, he also found significant deviations. It is possible that correlations will be developed based on oil composition as more information becomes available. While not eliminating the need for viscosity rneasurements, the method presented herein requires that only one viscosity measurement be made. The API gravity must also be known. The theory is based on the fact that the viscosity of paraffins (high gravity) changes less with temperature than does the viscosity of naph-thenes or aromatics (low gravity). The gravity. therefore, is used as a parameter to determine the slope of a straight line on the ASTM standard viscosity-temperature charts. The correlation is applicable only to Newtonian oils, and deviations due to thermal decomposition and nonhomo-geneity cannot be predicted. Oils containing additives have not been evaluated. PROCEDURE Fifteen oils were used in developing the correlation; eight were crudes and seven were processed oils. Oil gravities varied from 9.9" API (naphthene base) to 32.7&apos; API (paraffin base). The temperature range studied was 81 to 516F. Each oil used had a minimum of three viscosity measurements and each plotted essentially as a straight line on the ASTM charts. In all, 91 viscosity measurements were used in the correlation. Saybolt, rolling ball and capillary tube viscometers were used for the measurements. Viscosity data for Samples 1, 2, 4, 7, 10, 11 and 14 were obtained in Texaco, Inc. laboratories. The data for Samples 3, 5, 6, 8, 9, 12 and 15 were from Fortsch and Wilson,3 and data for Sample 13 were from Dean and Lane.&apos; All data points used in the correlation are plotted in Fig. 1. It is seen that some of the viscosity values deviated slightly from the straight-line plots at the higher temperatures. Properties of the oils after exposure to the

Jan 1, 1967

By S. F. Shaw

IT IS generally recognized that the natural gas absorbed in petroleum plays the leading r6le in moving the oil through the sands to the well and supplies the energy that delivers the oil to the surface when the wells are flowing. It is the purpose of this paper to note the amount of oil that could be lifted by this gas if 100 per cent. efficiency were obtained and the efficiencies actually obtained in the use of this gas. An article in MINING AND METALLURGY,&apos; stated that in a Wyoming field, with a pressure of 300 lb. per sq. in., there would be absorbed 65 cu. ft. of gas per barrel, or 1.55 cu. ft. per gal. of oil. The writer does not know if data have been obtained, by experiment, for pressures beyond those indicated in this article, therefore, for the purpose of illustration, he has projected this curve to obtain figures for pressures greater than 300 lb., although probably experiment will show that the extended line is not quite correct.

Jan 1, 1926

By N. P. Peterson

The Castle Dome copper deposit is of the porphyry type and occurs in a body of quartz monzonite intruded into the pre-Cambrian formations and possibly into the lower part of Paleozoic limestones. The hypogene sulphide minerals include pyrite, chalcopyrite, and a little molybdenite, sphalerite, and galena. Pyrite and chalcopyrite are disseminated or occur in small closely spaced veins. The hypogene mineralization shows distinct zoning in the distribution of pyrite and chalcopyrite, so that the mineralized area can be subdivided into pyritc and chalcopyrite zones. Supergene enrichment, though not extensive, played an important part in the formation of the ore body, the chalcopyritc zone being the pro-tore. Several thin, gently dipping, diabase sills occur in the mineralized area, and the copper content of the rock in and near these sills is substantially higher than average. The formation of the ore body was caused by the combined effects of three forces: I. Zoning of the hypogene mineralization, 2. Richer copper metallization associated with the diabase sills, and 3. Supergene enrichment. Introduction The Castle Dome copper deposit is about 5 miles west of Miami, Arizona, and about 3 miles west of the Miami-Inspiration ore body. The Castle Dome open-pit mine is on the south flank of Porphyry Mountain, a prominent landmark of the region. In 1941 the deposit was being explored by the Miami Copper Co., but the exploration program was cut short when the decision was made to develop the property as a war project. The Castle Dome Copper Co., Inc., was organized as a wholly owned subsidiary of the Miami Copper Co., and arrangements were made with Defense Plant Corp. for developing and equipping the mine and for construction of a concentrating plant. Construction and development work was begun in January 1942, and milling operations started June so, 1943. Since that date, the mine has been in continuous operation by the Castle Dome Copper Co, under a lease agreement with Defense Plant Corp. producing copper at an average rate of 4,000,000 lb per month. To January I, 1946 the total production amounted to 121,088,000 lb of copper,4,271 oz of gold, and 211,755 oz of silver. A detailed study of the copper deposit and the surrounding area was undertaken by the U. S. Geological Survey in June 1943. Other members of the Survey participating in the work were C. M. Gilbert, G. L. Quick, and J. P. Albers. The results of this work will be published as a comprehensive report by the Geological Survey. General Geology The Globe quadrangle, which includes the Castle Dome area, was studied by F. L. Ransome1 of the Geological Survey in 1901 and 1902. His maps and descriptions were

Jan 1, 1949

By N. P. Peterson

The Castle Dome copper deposit is of the porphyry type and occurs in a body of quartz monzonite intruded into the pre-Cambrian formations and possibly into the lower part of Paleozoic limestones. The hypogene sulphide minerals include pyrite, chalcopyrite, and a little molybdenite, sphalerite, and galena. Pyrite and chalcopyrite are disseminated or occur in small closely spaced veins. The hypogene mineralization shows distinct zoning in the distribution of pyrite and chalcopyrite, so that the mineralized area can be subdivided into pyritc and chalcopyrite zones. Supergene enrichment, though not extensive, played an important part in the formation of the ore body, the chalcopyritc zone being the pro-tore. Several thin, gently dipping, diabase sills occur in the mineralized area, and the copper content of the rock in and near these sills is substantially higher than average. The formation of the ore body was caused by the combined effects of three forces: I. Zoning of the hypogene mineralization, 2. Richer copper metallization associated with the diabase sills, and 3. Supergene enrichment. Introduction The Castle Dome copper deposit is about 5 miles west of Miami, Arizona, and about 3 miles west of the Miami-Inspiration ore body. The Castle Dome open-pit mine is on the south flank of Porphyry Mountain, a prominent landmark of the region. In 1941 the deposit was being explored by the Miami Copper Co., but the exploration program was cut short when the decision was made to develop the property as a war project. The Castle Dome Copper Co., Inc., was organized as a wholly owned subsidiary of the Miami Copper Co., and arrangements were made with Defense Plant Corp. for developing and equipping the mine and for construction of a concentrating plant. Construction and development work was begun in January 1942, and milling operations started June so, 1943. Since that date, the mine has been in continuous operation by the Castle Dome Copper Co, under a lease agreement with Defense Plant Corp. producing copper at an average rate of 4,000,000 lb per month. To January I, 1946 the total production amounted to 121,088,000 lb of copper,4,271 oz of gold, and 211,755 oz of silver. A detailed study of the copper deposit and the surrounding area was undertaken by the U. S. Geological Survey in June 1943. Other members of the Survey participating in the work were C. M. Gilbert, G. L. Quick, and J. P. Albers. The results of this work will be published as a comprehensive report by the Geological Survey. General Geology The Globe quadrangle, which includes the Castle Dome area, was studied by F. L. Ransome1 of the Geological Survey in 1901 and 1902. His maps and descriptions were

Jan 1, 1949

By K. K. Rao, M. E. Nicholson

IN a recent paper on the solubility of carbon in Ni-Cu alloys,&apos; Nicholson reported that the carbon solubility appeared to be limited by the electronic structure of the alloys and that the solubility was propor- tional to the density of states at the Fermi surface of the "3d" band. In order to determine whether or not this relationship is a general one, the study has been extended to other systems. One of these is the CO-Ni system. This system was chosen for study because 1) it is adjacent to Ni-Cu system in the periodic table, 2) the alloys of cobalt and nickel do not form stable carbides at 1000°C, and 3) the system exhibits no pronounced change (less than 1 pct) in lattice parameter from cobalt to nickel,&apos; see Fig. 1. The alloys were prepared from -200-mesh 99.92 pct Ni furnished by the International Nickel Co. and spectroscopically pure -200-mesh cobalt sponge obtained from Johns on-Matthey and Co. The powders were weighed, mixed, and pressed using a pressure of 40,000 psi. The samples were sintered for 24 hr at 900° C in dry hydrogen and then cold rolled to a thickness of 0.010 in. (Samples containing more than 70 wt pct Co had to be annealed after cold reductions of about 20 pct to avoid cracking.) The alloys were carburized at 1000°C in an alun-durn boat containing Acheson AUC powdered graphite in an atmosphere of carbon monoxide cylinder gas which had been passed through Ascarite to remove carbon dioxide. After carburizing for either 12 or 24 hr the samples were moved to a cool zone of the furnace to cool. Cooling required about 5 min. The samples were filed to remove about 0.005 in. from the surface and then analyzed using a conventional combustion-gravimetric analytic method. The results of this investigation are shown in the lower portion of Fig. 1 and Tabulated in Table I. The nickel percentages for these alloys were determined from the weights of the powders from which the alloys were made. The solubility of carbon in pure cobalt at 1000°C (fcc) is 0.25 wt pct. This is somewhat less than the 0.33 wt pct measured by Boecker.3 From his data, it would appear Boecker&apos;s samples were supersaturated before the equilibrating treatment and that a 20-min treatment was not long enough to achieve equilibrium. Since in the present investigation the 12 and 24 treatments produced essentially equal carbon contents, our figure is presumed to represent more nearly the equilibrium value. The solubility of carbon in nickel of 0.19 wt pct agrees quite well with the value of 0.21 wt pct

Jan 1, 1963

By W. R. Crane

THE support of roof in mines is dependent largely on the character of the top rock and its occurrence. The formations overlying the orebed in the Birmingham district are sandstone and slate. The sandstone occurs in beds of sufficient thickness to constitute important elements in the system of support; the slate is relatively strong but is thinly stratified. The alternating beds of sandstone and slate furnish an excellent combination in that the former are strong and the latter are impervious; however, their continuity is broken by jointing or slip planes. The thickness, and consequent weight of the overlying formations, is important in so far as the size and arrangement of pillars are concerned but may be relatively unimportant, temporarily at least, as a factor in the support of the roof. The position of the respective beds of sandstone and slate, their relative thickness, and their inherent weakness because of the occurrence of faults and folds, may exert a predominating influence on their support in place, which is affected only in part by the weight of the cover. The only mine in this district that operates under a cover of approximately 2000 ft. (1900 ft.) is the Shannon so-called "twin slope," situated 14,000 ft. southeast of the portal of the No. 7 mine of the Tennessee Coal, Iron, & Railroad Co. The conditions in this mine indicate what may be expected in other mines when the same depth is attained; in fact the effect of pressure is shown in several mines that have not reached two-thirds of that depth. It is evident then that conditions affecting the support of roof will not improve but rather will become more difficult with the extension of the mines into the valley under a constantly increasing weight of cover. Further, the difficulties will probably be augmented by the occurrence of faults and other disturbed ground, the presence and position of which at present are largely unknown.

Jan 9, 1924

T.B.King (Depaytment of Metallurgy, Massachusetts Institute of Technology)— A valuable contribution of the authors is in the factual information which they have been able to gather; this type of information is quite difficult to obtain. In many respects, however, it would have been better if they had not subsequently embarked on a discussion of the chemistry of the converter process. It seems inconceivable that the authors do not refer to the papers of Schuhmann and his associates14 which have set the thermodynamic foundation for the whole copper smelting operation. In addition, a very useful review on the physical chemistry of copper smelting by Ruddle 5 appeared as long ago as 1953. An examination of this literature would have convinced the authors that there is no cause to be surprised at a correlation between the magnetic content and the silica content of converter slags, though they rightly point out that one should distinguish between the total magnetite content of the slag and the amount of magnetite which may be considered to be in solution. It is not true that the lowest melting converter slag is that corresponding to the eutectic between ferrous oxide and silica. The simplest slag system which can be considered is a three-component system, since both ferric and ferrous iron are present. As Schuhmann, Powell, and Michal have shown, there are lower melting compositions than this eutectic in the ternary system. The most unfortunate impression given by this paper is that the driving force for chemical reaction is determined by the heat of reaction: of course the entropy change must be taken into account. It would have been more correct to list, in Table VI, values for free energies of formation. Nor can it be said that the data in Table VI represent the "best available data." They do not corregtond with any of the recent, acknowledged sources. F. E. Lathe and L. Hodnett(author&apos;s reply)— We are pleased that Dr. King finds the factual information in our uawer of some interest. Dr. King suggests that it would have been better if our analysis and discussion of the data had been omitted, largely because our list of references is so. incomplete. If he will carefully read our introduction, he will see that the questionnaire was sent out in the hope of obtaining data which would throw light on certain questions relating to the use of converter refractories. We did not attempt (nor would the AIME have published!) a complete review of the literature on copper converting, as Dr. King has apparently assumed, nor indeed a complete analysis of the data submitted, but tried only to find a sound basis for the choice of refractories, taking into consideration common variations in converter practice. We hope our paper indicates that, by raising the silica content of the converter slag and operating at a higher temperature, the normal circulating load of magnetite can be greatly reduced, and the whole reverberatory-converter operation improved to a major degree, with resultant important savings. Under such operating conditions, chrome-magnesite brick may be expected to stand up better than those of straight magnesite. Regardless of the choice as between these brick types, however, we find the cost of converter refractories to be so low in comparison with other converter costs as to justify operation under the more severe conditions suggested. Valuable as are the papers by Schuhmann and associates and the book by Ruddle, we make no apology for omitting reference to them, nor for using heats of reaction without mention of entropy changes or free energies of formation. Our primary object was to interest the practicing copper metallurgist, with whose language we may claim to be fairly familiar; we think it would have been unwise to include the highly theoretical phases of the subject which Dr. King suggests. The interest shown in our preliminary paper presented at the New York meeting in 1956, and the trends in practice whtat we have observed since that time, suggest that we did not wholly miss the target. In conclusion, we sincerely hope that Professors King and Schuhmann will independently review the data obtained in our questionnaire and submit a paper giving their own recommendations as to the choice of refractories and the particular converter operating conditions which will result in the lowest overall cost of copper smelting and converting.

Jan 1, 1960

By Paul Averitt

The U. S. Geological Survey has been actively engaged in work on coal for more than 50 years. During this long period we have released more than 300 publications containing information about coal and coal reserves, including details on the thickness, distribution, and quality of coal, and the structure and stratigraphy of the coal-bearing rocks. We have also cooperated with state surveys on many similar publications. For many areas these reports are the only reliable published data on our resources of coal. During the last several years we have substantially increased our activities above the wartime low. We now have field work in progress, or reports in preparation, on 13 detailed mapping projects distributed as follows: 1. Pennsylvania anthracite field 2. Casselrnan Basin, Md. 3. Leslie County, Ky. 4. Deep River field, N. C. 5. Coosa field, Ala. 6. Powder River field, Mont. 7. Spotted Horse field, Wyo. 8. Yampa field, Colo. 9. Paonia field, Colo. 10. Trinidad field, Colo. 11. Durango field, Colo.

Jan 1, 1949

By S. Ramachandran, R. G. Woolery

INTRODUCTION Union Carbide is currently operating an in-situ leach project on the Palangana Dome area in Duval county. This deposit meets all the requirements for in-situ leach in that the ore (1) is below the water table, (2) is in a permeable horizon, (3) is amenable to chemical leaching, and (4) is confined by impervious layers. This project has been under commercial production since 1976, and its capacity has been expanded on three occasions since going on-stream. Recently, additional uranium reserves were discovered on the Rogers-Cardenas (R-C) property about 32 km north of the Palangana operation. The ore is located within the Oakville sands and its characteristics are quite similar to those of the ore at Palangana. Both are an unconsolidated Arkosic sand high in clay and calcium carbonate. The R-C ore, however, is somewhat coarser with a mean particle size of 0.15 mm as compared to a mean particle size of 0.07 mm for the Palangana ore. In all respects it would appear that this ore would be a candidate for in-situ leach as a satellite operation to Palangana. Unfortunately, R-C ore is above the water table and, therefore, not amenable to the Palangana practice. Because of the limited known reserves in this deposit, it is readily apparent that conventional mining and milling are out of the question. However, because of its proximity to our Palangana operation, it seemed worthwhile to consider other options. The most viable route based on our past experience was to heap leach the ore. Our recent success at our Gas Hills facility and our Maybell operation, in employing a heap leach practice to our marginal reserves seemed to be a logical approach for processing this ore. Our experiences at both locations are described in "Heap Leaching - A Case History" by R. G. Woolery et al., Mining Engineering, March 1978. In both instances the process is an acid leach circuit and acid consumption averages 20 kg/t H2SO4. A preliminary feasibility study showed that because of the high strip ratio required for the R-C project to be successful, additional ore reserves must be located and that a method of heap leaching with an alkaline circuit would have to be developed. As a result of this paper study, the decision was made to proceed with a program of additional exploration drilling to determine the total ore reserves that could be mined economically. The Mining Department will evaluate each ore zone for cutoff grade, strip ratio, and expected mining cost. At the same time, a laboratory program to evaluate the available core samples for amenability to heap leaching with respect to an estimate of uranium recovery and processing costs was developed. This program is currently in progress, and at this time, we are just completing our process amenability study. BENCH-SCALE EVALUATION OF THE R-C ORE The initial bench-scale slurry leach tests on the R-C ore showed an acid consumption in excess of 200 kg/t H2SO4. These data, of course, discouraged us from considering this process route. Not only would the acid cost be prohibitive, but the gypsum generated by the reaction of the sulfuric acid with the calcium carbonate of the ore would severely effect the percolation of the lixivant. For this reason, the laboratory program was directed toward an alkaline circuit compatible with heap leaching. Because of the proximity of the R-C property to our Palangana operation, it seemed advisable to integrate the processing of this ore into the production at Palangana. Doing so would enable us to bring the R-C property into production by merely enlarging our present facilities at Palangana; otherwise, construction of a grass roots plant would be necessary. Ideally, the simplest method would be to construct the heaps at Palangana and employ an ammonium carbonate/bicarbonate leachant compatible with the in-situ production liquor. The product liquors could then be co-mingled or processed separately as desired. To determine if this goal was practical, samples of the R-C ore were obtained, and a laboratory program initiated. Heap leach amenability testing consisted of preliminary bench-scale evaluation to determine optimum solution strength and ultimate uranium recovery, followed by small column tests to confirm the bench-scale metallurgy and to determine percolation characteristics. These bench-scale tests are being followed by pilot-scale testing approximating field conditions. As expected, the bench-scale tests showed that the dissolution rate is considerably slower for alkaline leach than has been our experience in acid leaching. Because of the slower reaction rates, product liquor grades will be lower than for acid, as greater volumes of solution are required for satisfactory uranium extractions. The greatest influence on reaction times found in the laboratory was the carbonate/bicarbonate strength and oxidant addition. However, the higher salt concentration reduced the efficiency of the IX resin circuit and about 25g/L salt proved to an upper limit compatible with subsequent IX treatment. The oxidant contributed significantly to the early extraction rate but seemed to have only minimal effect on the total practical U308 extracted or the time required to achieve it. This variable will require larger scale testing to determine if the added cost of the oxidant is actually justifiable. Thus, the small-scale laboratory slurry tests, based on the 0.088% U308 sample available, indicate that leaching at 25g/L ammonium carbonate/bicarbonate, with or without oxidant, we might expect an 80-85% U308 extraction on this ore.

Jan 1, 1979

By L. J. Buck

Lead smelting in the ore hearth, first practiced in Europe, whence the traditional name "Scotch hearth, and later developed in America to its present efficiency, is based upon the roast reduction process and is applicable to relatively high-grade galena concentrates. The roast reduction process embraces not only the roast reactions: 2PbS + 70 = PbO + PbSO4, + SO2 PbS + 2PbO = 3Pb + SO2 PbS + 2PbSO4 = Pb + 2PbO + 3SO2 but also, by reason of the addition of extraneous fuel, the reduction reaction: 2PbO + C = 2Pb + CO2 Thus in the ore hearth the roasting of lead sulphide, forming lead oxide and lead sulphate, the reaction of lead oxide and lead sulphate with lead sulphide producing lead, and the reduction of lead oxide by carbon producing lead, go on simultaneously. The Ore Hearth Function of the Ore Hearth.—The ore hearth, through the roast reduction process, provides for the extraction of lead and the immediate disposal of the smelting products—fume, hearth slag and lead, all of which are subsequently treated as follows: the fume is recovered in the flue system and baghouses, mixed with concentrates, and returned to the ore hearth in the form of ore mix; the hearth slag is smelted in a blast furnacc for the further recovery of lead; and the lead is refined and cast into a marketable product. Concentrates Amenable to Ore-hearth Smelting.—Only concentrates that are high in lead content, 70 per cent or more, are ordinarily considered amenable to ore-hearth smelting. Silica should not exceed 2 per cent and iron is allowable only to 4 per cent. These slag-forming constituents occlude both sulphide and mctallic lead, prevent the elimination of sulphur and decrease the lead extraction. Lime and magnesia are beneficial in that both favor a dry slag, promote elimination of sulphur, and prevent occlusions of metallic lead.

Jan 1, 1937