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By A. A. Smith

THE application of indium to bearings was first advocated by C. F. Smart1 in 1938, for the purpose of inhibiting corrosion of certain bearing alloys. He found that a thin layer of indium plated on the bearing surface and subsequently diffused at elevated temperatures markedly decreased the rate of corrosion in lubricating oils containing organic acids. Most of his work was done on cadmium-base alloys, although he showed that the treatment was also effective for copper-lead and other alloys. Since their introduction, indium-treated bearings have found considerable use, particularly in connection with airplane motors.2 Undoubtedly the effectiveness of the indium plate in resisting corrosion depends upon its relative concentration on the bearing surface after the diffusion treatment and after prolonged use at operating temperatures. Some diffusion experiments were made on a number of base metals that had been indium plated, to determine the depth of penetration after heating for reasonably long periods of time. As only the practical aspects were desired, no attempt was made to determine actual diffusion constants. METHOD Two different series of experiments were carried out to study the diffusion of indium into several base metals. Cast cylinders about one-inch diameter were machined to 0.750-in. diameter, so as to remove all surface imperfections, and final finishing was done with oo sandpaper. The original diameters were very carefully measured with micrometers and then a layer of indium was electrolytically deposited on the cylinders from a sulphate bath.3 In the first series of experiments, about 0.002 in. of indium was deposited and diffusion was carried out at 150°C. for a period of 500 hr. in a thermostatically con- trolled electric oven. After the diffusion period, successive layers were machined from the specimens and analyzed spectro- graphically and chemically-the spectro- graphic examination serving as a guide to determine which layers should be chemically analyzed. Apparently there was some surface oxidation of the indium plate due to the long heat-treatment, which may have had some effect on the diffusion. Also, the cadmium alloy showed some incipient melting, which probably was caused by the formation of a low-melting eutectic. To eliminate these factors, another series of experiments was carried out by heating in an electrically heated, automatically controlled, oil bath at lower temperature. The preparation of the specimens was the same as before except that a layer of indium 0.0001 in. thick, an amount commonly used commercially on bearings, was deposited on each of the cylinders. Diffusion was carried out at l20°C. for a period of 200 hours.

Jan 1, 1944

By Felix B. Shay

The Agua Fria mine is in eastern Honduras at Lat. 140" 06' N., Long. 86" 36' W. Danli, the seat of the local municipal government, is the nearest town. The property is approximately 44 miles east of Tegucigalpa, the terminus of a truck road from the Pacific coast. The mine lies at an altitude of 3300 ft. with surrounding elevations as high as 6000 feet. For 20 years the oxidized croppings of the Agua Fria vein had been worked intermittently when in 1927 an attempt was made to open the property. A shaft was sunk 300 ft. and considerable ore was blocked out, but before a mill could be installed financial difficulties arose and the mine reverted to the original owners. Erle Y. Halliburton, of Los Angeles, purchased the property in 1934. At that time a careful consideration of all factors involved showed that air transport was the only practical method of moving the equipment and machinery necessary to commence production. The other alternative was by muleback, which required sectionalizing all equipment to a maximum load of 200 lb., and during the rainy season this type of freighting was subject to serious delays. It was decided to purchase a trimotor Ford powered by three Pratt and Whitney Wasp B motors developing 400 hp. each. There were several reasons for the choice: (1) this type of plane was available at a reasonable price, owing to its replacement by faster models on most of the airlines in the United States; (2) the ship was capable of carrying a pay load of 3000 Ib. over the required distance; (3) since this airplane had been widely used there would be no difficulty in securing competent personnel or in purchasing spare parts. It was necessary to rebuild the plane for cargo servicc. The windows were removed and the fuselage was entirely covered. A hatch was cut in the top of the fuselage to receive heavy crane loads and pieces of freight that were too large or too long to go through the rear door. The seats were removed and the interior of the plane was lined with Duralumin sheets. A heavy-duty landing gear was installed and the floor was reinforced. The nearest site for an airport was two miles north of the mine. The necessary grading was done by a small track-type tractor with a wheeled grader, which was brought in overland. Six weeks was required for bring-

Jan 1, 1940

By Felix B. Shay

The Agua Fria mine is in eastern Honduras at Lat. 140" 06' N., Long. 86" 36' W. Danli, the seat of the local municipal government, is the nearest town. The property is approximately 44 miles east of Tegucigalpa, the terminus of a truck road from the Pacific coast. The mine lies at an altitude of 3300 ft. with surrounding elevations as high as 6000 feet. For 20 years the oxidized croppings of the Agua Fria vein had been worked intermittently when in 1927 an attempt was made to open the property. A shaft was sunk 300 ft. and considerable ore was blocked out, but before a mill could be installed financial difficulties arose and the mine reverted to the original owners. Erle Y. Halliburton, of Los Angeles, purchased the property in 1934. At that time a careful consideration of all factors involved showed that air transport was the only practical method of moving the equipment and machinery necessary to commence production. The other alternative was by muleback, which required sectionalizing all equipment to a maximum load of 200 lb., and during the rainy season this type of freighting was subject to serious delays. It was decided to purchase a trimotor Ford powered by three Pratt and Whitney Wasp B motors developing 400 hp. each. There were several reasons for the choice: (1) this type of plane was available at a reasonable price, owing to its replacement by faster models on most of the airlines in the United States; (2) the ship was capable of carrying a pay load of 3000 Ib. over the required distance; (3) since this airplane had been widely used there would be no difficulty in securing competent personnel or in purchasing spare parts. It was necessary to rebuild the plane for cargo servicc. The windows were removed and the fuselage was entirely covered. A hatch was cut in the top of the fuselage to receive heavy crane loads and pieces of freight that were too large or too long to go through the rear door. The seats were removed and the interior of the plane was lined with Duralumin sheets. A heavy-duty landing gear was installed and the floor was reinforced. The nearest site for an airport was two miles north of the mine. The necessary grading was done by a small track-type tractor with a wheeled grader, which was brought in overland. Six weeks was required for bring-

Jan 1, 1940

By Leonard D. Jaffe

IN recent years, the advantages of tempered martensite as a microstructure for steel parts have been well established. For parts that must not fracture brittlely when loaded at high rates, at low temperatures, or under conditions of combined stress, tempered martensite is especially desirable. One of the factors that determine whether a steel part will have a martensitic structure after heat treatment is the hardenability of the steel. Despite the large amount of recent research, there are still many unsettled basic questions concerning hardenability and its relation to chemical composition, and, indeed, considerable disagreement among investigators in the field. Much of the uncertainty that beclouds the existing knowledge of hardenability arises from the fact that hardenability is still a wholly empirical concept, and has not been quantitatively related to the more fundamental knowledge of isothermal decomposition of austenite. Hardenability is, essentially, a measure of the time required for small percentages of austenite to decompose, during continuous cooling, to pearlite, bainite, proeutectoid ferrite, or proeutectoid carbide. In the practical heat-treatment of steel, also, the decomposition of austenite is usually not isothermal but rather during cooling, or, in general, is not at a single constant temperature. Thus, both hardenability and practical heat treatment involve anisothermal decomposition of austenite. In a previous publication1 the author and his collaborators outlined the principles that govern the relations between isothermal and anisothermal decomposition of austenite, and reviewed the experimental information available in 1945. To avoid repetition, familiarity with that article (Paper 1) will be assumed, and attention given primarily to work carried out or published since 1945. The method generally employed was explained in Paper 1: anisothermal transformation is studied through investigation of the effect of holding austenite at one temperature upon its subsequent decomposition at another temperature. Both temperatures are below the stability range of the austenite. This paper is confined to studies in which only a small amount of transformation occurred at the first temperature. This is the only case of importance in practical heat treatment when an essentially martensitic structure is sought. MATERIALS AND PROCEDURE It was necessary to use different steels for the studies involving proeutectoid ferrite and for those involving proeutectoid carbide, because the two constituents have not been found to form in the same steel. Several steels were tried and rejected

Jan 1, 1947

By R. R. Knuckey

Having been associated with the operation of the de Vooys process for coal, which has treated 13,000,000 tons per annum, and recognizing the process as of value in ore sorting, Huntington, Heberlein and Co. Ltd. (Great Britain) in 1936 made an examination of all other existing methods. In none of the pioneer applications had recognition been given to the necessity for maintaining certain conditions that are essential for sorting certain ores if a truly efficient separation by gravity alone is to be assured. Consequently, in 1936-1937 Huntington, Heberlein and Co. Ltd., with Pearson, employing certain ideas connected with the de Vooys application, commenced the development of a separator that would allow conformance with optimum conditions in the zone where separation is effected. Later, Holt and Wade added to the value of the original application by Wuensch, Rakowsky and others, whose conceptions are being successfully operated today. Meantime, Trostler had enhanced the value of the original H-H. process by improvements that are now embodied in operating plants, and today the process incorporates an improved apparatus and a stable medium that produce the optimum conditions for separation. These conditions are: 1. Elimination of disturbing currents in the separating zone; 2. Reduction to a minimum of medium-density differences throughout the separating zone; 3. Stable medium of minimum viscosity to permit the unretarded fall of mineralized particles; 4. Removal of products of separation with minimum disturbance to the separating zone. Design of Separator The H-H. sink-and-float separator consists of a pyramidal tank in which the separation actually takes place, connected at the bottom with an inclined bucket elevator. The tank and the elevator are filled with medium to within a few inches of the top of the tank. Medium is fed into the separator at a variety of points, according to the type of ore being treated, and it is withdrawn from the separator at two points: (1) together with the float product, by means of a variable-speed addle, at the surface of the medium, and (2) by flowing over an adjustable weir cut in the side of the elevator casing. The prepared ore is fed onto the surface of the medium. The intermediate paddles, which revolve at a slow speed of 2 to 5 r.p.m., move the floating ore across the bath from feed end to float-discharge end. The ore, which is heavier than the medium, sinks below the sweep of the addles and reaches the boot of the bucket elevator, which discharges it without disturbance of the separating zone. These arrangements for withdrawing medium from the separator allow the use of minimum flows, which connote minimum

Jan 1, 1947

By R. R. Knuckey

Having been associated with the operation of the de Vooys process for coal, which has treated 13,000,000 tons per annum, and recognizing the process as of value in ore sorting, Huntington, Heberlein and Co. Ltd. (Great Britain) in 1936 made an examination of all other existing methods. In none of the pioneer applications had recognition been given to the necessity for maintaining certain conditions that are essential for sorting certain ores if a truly efficient separation by gravity alone is to be assured. Consequently, in 1936-1937 Huntington, Heberlein and Co. Ltd., with Pearson, employing certain ideas connected with the de Vooys application, commenced the development of a separator that would allow conformance with optimum conditions in the zone where separation is effected. Later, Holt and Wade added to the value of the original application by Wuensch, Rakowsky and others, whose conceptions are being successfully operated today. Meantime, Trostler had enhanced the value of the original H-H. process by improvements that are now embodied in operating plants, and today the process incorporates an improved apparatus and a stable medium that produce the optimum conditions for separation. These conditions are: 1. Elimination of disturbing currents in the separating zone; 2. Reduction to a minimum of medium-density differences throughout the separating zone; 3. Stable medium of minimum viscosity to permit the unretarded fall of mineralized particles; 4. Removal of products of separation with minimum disturbance to the separating zone. Design of Separator The H-H. sink-and-float separator consists of a pyramidal tank in which the separation actually takes place, connected at the bottom with an inclined bucket elevator. The tank and the elevator are filled with medium to within a few inches of the top of the tank. Medium is fed into the separator at a variety of points, according to the type of ore being treated, and it is withdrawn from the separator at two points: (1) together with the float product, by means of a variable-speed addle, at the surface of the medium, and (2) by flowing over an adjustable weir cut in the side of the elevator casing. The prepared ore is fed onto the surface of the medium. The intermediate paddles, which revolve at a slow speed of 2 to 5 r.p.m., move the floating ore across the bath from feed end to float-discharge end. The ore, which is heavier than the medium, sinks below the sweep of the addles and reaches the boot of the bucket elevator, which discharges it without disturbance of the separating zone. These arrangements for withdrawing medium from the separator allow the use of minimum flows, which connote minimum

Jan 1, 1947

By M. W. Dumais, E. P. Abrahamson

In a paper by Abrahamson and Alexander concerning a series of correlations between electronic structure and property changes, it was hypothesized that each solute atom in a dilute solution had a characteristic volume of solvent atoms over which some form of electronic interaction occurred. From this description it is possible to picture the solute atoms as occurring in an ordered array for the most efficient arrangement. A recent review article2 has shown that some form of anomaly exists for a whole series of properties when dilute additions of tungsten are made to iron, of. Fig. 1. Any mechanism to fit one property must also be amenable to mechanisms for the other properties. The possibility of some structural change in this region appears to offer the best hope. The system chosen for investigation was the iron tungsten system. Zone-refined iron with zone-leveled tungsten additions were supplied by Materials Research Corp. as 3/8-in.-diam bars. Transmission electron microscopy was used to study the structure. One-quarter inch specimens of each alloy were cold-rolled to 10-mil sheets, cleaned, punched to electron-microscope specimen-holder size, and annealed in vacuum for 2 hr at 800°C and 2 hr at 700°C followed by a furnace cool. The specimens were attached to a platinum wire and the specimen edge and wire were painted prior to thinning in a solution of 10 pct perchloric acid in glacial acetic acid. It was believed that if ordering were present due to electronic interactions the polishing procedure chosen would provide a simultaneous combination of chemical attack and electropolishing to exhibit the ordered structure. The chemical attack would be expected to penetrate those areas equidistant from the tungsten atoms, while the electropolishing would serve to "plane off" the volumes where the chemical attack had taken place. If one of the ordered tungsten planes were parallel to the specimen face at the edge of the thinned hole, a plate or micalike structure would be expected. A typical result of this polishing is shown in Fig. 2. Here a series of plates can be seen lying one upon the other. In several other specimens the plates parallel to the electron beam have been observed, behaving like slices of bread in a loaf. These edge views coupled with moire patterns indicate that the minimum plate thickness is 37 to 42A. This is equivalent to a concentration of 455 at. ppm of W if a simple cubic superlat-tice exists, or one tungsten atom every twelve iron atoms on the ordered plane. It was further noted that the base iron exhibited no signs of platelike structure. As the concentrations increased above 100 at. ppm W the amount of plate material increased. Among the specimens investigated was that shown

Jan 1, 1967

By Frank R. Milliken, Robert Goodwin

In September 1937, the Ohio Copper Co. inaugurated the treatment of its copper-bearing mill tailings at Lark, Utah. These tailings had been accumulated during the regular operation of the Ohio Copper mine and mill over the period 1907 to 1919. Gravity concentration had been employed except for a comparatively short period in 1919, when flotation was used. The ore milled during that period assayed about I per cent Cu. The chief copper mineral, chalcocite, was disseminated in a shattered quartzite, mainly along fracture or cleavage planes. The quartzite gangue was extremely hard and grinding to liberation resulted in excessive sliming of the "sooty" chalcocite, with consequent high tailings losses. Recovery on the seven million tons treated was probably less than 60 per cent of the total copper content. As early as 1918, attempts had been made to sample the dump. In 1928, a complete sampling program was undertaken. The sampling consisted of seventy-four 5-in. post auger holes drilled to the full depth of the dump at regular intervals of 250 ft. Each 10-ft. portion thereof comprised a sample for analysis. Each sample was screened, rolled, divided into three parts, dried and assayed. The reject was retained for testing purposes. From the depth of the auger holes a contour map of the dump was plotted, Fig. I, which gives percentages of tonnage and copper content. The tonnage within the limit of the drill holes was estimated at about five million tons with an average content of 0.42 per cent Cu, 31 per cent of which (0.13 per cent) was acid-soluble. A screen analysis of the composite dump sample is shown in Table I. Flotation tests by several companies and individuals were then made, the results of which ranged from 62 to 74 per cent recovery. On the basis of these, a 200-ton pilot plant employing straight flotation was built and operated for a few months. Operation of this plant was unsuccessful, as no provision had been made to handle the copper-bearing solutions resulting from the water-soluble copper, a point that was overlooked in the laboratory test work. Attempts to decant the copper-bearing solution for precipitation in the usual manner ahead of flotation were not satisfactory because of low water recovery (owing to poor settling characteristics of the pulp) and contamination of the copper precipitate with unsettled slime. Low copper prices in 1931 necessitated the termination of testing. With the stimulus of higher copper prices in 1937, a review was made of metallurgical

Jan 1, 1943

By R. M. Asimow

The magnetization curves for specimezs containing superparmagnetic particles are considered. It is shown that the curves may differ from a Langezlin function because of particle anisotropy, particle interactions , and the presence of a size distribution of particles. The various causes of particle anisotropy are considered, and magnetization curves for specimens with cubic and uniaxial anisotropy are presented. The magnetization curves depend on the In the particle-size below 100A it is quite difficult to obtain quantitative size and shape information on precipitate particles. All of the available techniques have their limitations; however, quantitative applications of superparmagnetism have not been developed as fully as possible due to a lack of analysis of the complete magnetization curve. With such an analysis it appears possible to obtain from a single magnetization curve of a specimen containing randomly oriented SPM (superparamagnetic) particles a quantitative volume distribution and a description of the particle shape.* sign of the anisotropy energy and in the case of a negative uniaxial anisotropy (prolate spheroids) the resulting curves may differ very considerbly from a Langevin function. A graphical technique is developed whereby it is possible to separate out the effects of the size distribution, the particle interaction, and the particle anisotropy. Thus it is possible to obtain quantitative values for both the particle shape and the size distribution. Langevin function2 I/Is = ctnh(HVIs/k T)-k T/HVIS [ 1 ] where I, is the saturation magnetization, V the particle volume, H the applied field, T the particle temperature, k the Boltzmann constant, and I the magnetization in the direction of the applied field. Such behavior is rarely if ever observed in actual systems. One explanation for this is that the SPM particles do possess a significant anisotropy and thus the magnetic energy associated with the particle depends on the direction of magnetization in a more complicated way than for the isotropic particle of Eq. [I]. In general the magnetic energy E of an isolated SPM particle embedded in a nonmagnetic matrix can be expressed as a sum of several contributions: E = Eh + Ed+Es+Ec [2] where Eh is the external field energy, Ed the energy resulting from the particle demagnetizing factor, E, the magnetostriction energy, and E, the crystalline anisotropy energy. In each term we consider only the part dependent on the orientation of the magnetization vector and the particle axes as the shape of the magnetization curve will be determined by these contributions. Following the same procedure used in deriving the Langevin function,

Jan 1, 1965

By Robert Goodwin, Frank R. Milliken

In September 1937, the Ohio Copper Co. inaugurated the treatment of its copper-bearing mill tailings at Lark, Utah. These tailings had been accumulated during the regular operation of the Ohio Copper mine and mill over the period 1907 to 1919. Gravity concentration had been employed except for a comparatively short period in 1919, when flotation was used. The ore milled during that period assayed about I per cent Cu. The chief copper mineral, chalcocite, was disseminated in a shattered quartzite, mainly along fracture or cleavage planes. The quartzite gangue was extremely hard and grinding to liberation resulted in excessive sliming of the "sooty" chalcocite, with consequent high tailings losses. Recovery on the seven million tons treated was probably less than 60 per cent of the total copper content. As early as 1918, attempts had been made to sample the dump. In 1928, a complete sampling program was undertaken. The sampling consisted of seventy-four 5-in. post auger holes drilled to the full depth of the dump at regular intervals of 250 ft. Each 10-ft. portion thereof comprised a sample for analysis. Each sample was screened, rolled, divided into three parts, dried and assayed. The reject was retained for testing purposes. From the depth of the auger holes a contour map of the dump was plotted, Fig. I, which gives percentages of tonnage and copper content. The tonnage within the limit of the drill holes was estimated at about five million tons with an average content of 0.42 per cent Cu, 31 per cent of which (0.13 per cent) was acid-soluble. A screen analysis of the composite dump sample is shown in Table I. Flotation tests by several companies and individuals were then made, the results of which ranged from 62 to 74 per cent recovery. On the basis of these, a 200-ton pilot plant employing straight flotation was built and operated for a few months. Operation of this plant was unsuccessful, as no provision had been made to handle the copper-bearing solutions resulting from the water-soluble copper, a point that was overlooked in the laboratory test work. Attempts to decant the copper-bearing solution for precipitation in the usual manner ahead of flotation were not satisfactory because of low water recovery (owing to poor settling characteristics of the pulp) and contamination of the copper precipitate with unsettled slime. Low copper prices in 1931 necessitated the termination of testing. With the stimulus of higher copper prices in 1937, a review was made of metallurgical

Jan 1, 1943

By R. J. Mellen

IN a chloridizing roasting and cyanidation plant treating a manganese-silver ore, certain experiments pointed out the way to an improvement in roasting conditions by a change in the degree of crushing. The following short article is simply a description of the various changes that were necessary throughout the plant and of the results obtained. There is nothing unusual in these changes but they are of some interest in that they constitute a good many of the alterations usually resorted to in increasing production. In the beginning, the changes were brought about by an experiment that was not successful, but which clearly pointed to means of improving the operation, and is an example of the fact that often experimental work may result in failure and still be of value in arriving at correct fundamentals. This experiment also shows how far a metallurgist can get out of line in his deductions, particularly in a field with which he is not thoroughly familiar. The ore at the plant of Compania Minera Kildun y Anexas is crushed through a 13-in. gyratory, a 3-ft. standard Symons cone crusher and a set of 16 by 42in. rolls, the latter in closed circuit with two Tyler 400-mesh screens. The crushed ore is then conveyed to two fine-ore bins. These provide storage and feed for two 6 by 90-ft. kiln roasters, which discharge directly to coolers. The coolers in turn discharge to a 6 by 12-ft. ball mill, where the ore is ground for cyanidation. Laboratory and pilot-plant tests previous to the mill construction had indicated that minus ¼ -in. crushing was satisfactory for roasting, although results did not vary greatly with some coarser or considerably finer crushing. The plant was designed for a capacity of 300 tons per day on this basis. After a few months operation, the initial operating problems were overcome and the anticipated milling rate was reached but the results being obtained were not really good. The plant seemingly did not have much excess capacity in any department but the roasting operation was the most critically limiting factor both as to tonnage and the results obtained. If this could be improved there seemed some possibility of improving the total tonnage rate to some extent. Segregation of the calcine into various screen sizes showed that the coarser sizes ,were not giving the expected recoveries. In all of the laboratory and pilot-plant testing, any limiting conditions as to the percentage of salt used, temperatures, feed rates, etc., had affected the fine material more than it had the coarse, and even in making laboratory tests at this time the mill conditions could not be exactly duplicated. Further laboratory testing indicated that a segregation of the crushed ore into a coarse and fine product for roasting might result in some improvement in roast-

Jan 1, 1945

By D. B. Carson, D. L. Katz

Natural gases under pressure form crystalline hydrates with water. Experimental data are reported on four-phase equilibrium for the methane-propane-water, methanc-pentane-water, and methane-hexane-water systems. Temperatures and pressures for equilibrium between gas, water-rich liquid, hydrocarbon-rich liquid, and hydrate were measured, as well as the percentages of methane and propane in the hydrate. The data indicate that natural gas hydrate behaves as a solid solution and that pentanes and heavier hydrocarbons do not enter into the solid phase. Vapor-solid equilibrium constants are presented that permit the approximation of the conditions for hydrate formation, from the composition of a gas. Introduction Natural gas hydrates have been the object of considerable research in recent years, because of the trouble they have caused in the natural-gas and natural-gasoline industries. Natural gas hydrates are white crystalline compounds of water and gas, which, under pressure, exist at temperatures considerably above the freezing point of water. Because of the relatively high temperatures at which the hydrates exist, they become a nuisance in high-pressure gas operations where water is present, since their formation causes partial or complete plugging of valves and pipes. From a practical standpoint, the trouble incident to hydrate formation has been solved by dehydration of the gas before it enters the plant or pipe line, or by other remedial measures. Extensive work on gas hydrates in the latter part of the nineteenth century, by Villardl and others, gave data on methane hydrate and on ethane hydrate. Schroeder summarized these theoretical studies2 Hammerschmidt3 introduced the information on gas hydrates to the gas industry and Dcaton and Frost4-6 gathered a considerable number of data on the behavior of natural gas hydrates. Recent papers7,8 have extended both the data and the theory of these hydrates. At present, data are available on tem-peraturcs and pressures of hydrate forma. tion for pure methane, ethane, propane, n-butane, methane-ethane mixtures, meth-ane-propane mixtures, methane-butane mixtures, and some Is natural gases, all- in the Presence of excess water. However, up to the Present time no method has been available for predicting the temperature at given pressures or the pressure at given at at which natural gases will form hydrates if hydrates with water. 'This Paper Presents data on the methane-propane-water, methane-pentane-water and methane-hex-""e water systems in four phases, one of which is the hydrate. he application of the phase rule to the methane-~ro~ane-water data Proves that hydrates behave as solid solutions. The data of this paper and those in the literature were used to develop vapor-solid. equilibrium constants,8 which permit prediction of conditions of hydrate formation from the analysis of a natural gas.

Jan 1, 1942

By O. W. Bilharz

Acid mine-water drainage is a serious problem with many mines in the Tri-State zinc and lead mining district. Particularly is this true when large volumes must be considercd in unwatering old mines that have refilled with water which has dissolved the products of oxidation and has become chalybeate in character. The experiences discussed in this paper have to do specifically with the unwatering of a comparatively small area west of the town of Baxter Springs, Kans., where the pumping job was complicated by water of this type and where the extremely high sulphate content particularly effected that phase of mine drainage having to do with stream pollution. Location, Geology and History The Baxter Springs area of the Tri-State mining field is situated just west of the town of Baxter Springs and is a northeasterly extension of the main mineralized zones of the Oklahoma part of the district (Fig I). The mining area is traversed with well defined shear zones which provide easy circulation of ground waters and the ore bodies are found along these shear zones as localized enrichments wherever the structural conditions are favorable (Fig 2). The ore-bearing strata is Mississippian with an overlain capping of shale of Penn- sylvanian age. This shale strata carries much pyrite and marcasite at its base and undoubtedly was a contributing factor to the subsequent high iron content of the water. The area originally contained a sulphide water which discharged as an artesian flow in springs within the town and gave Baxter Springs its name. These springs were famed for many years for their mineral water. This original sulphide water was first removed about 1917 and active mining was carried on west of the town during subsequent high metal price periods. Many small mines were opened, with each localized enrichment becoming a separate mine. During the depression mining ceased and all mines in the west Baxter area were allowed to flood. No serious effort was made to again unwater until higher prices caused from the demand for metal for World War II once more focused attention 011 this part of the field. Original Mining Results in Oxidation As plans for unwatering progressed, it soon developed that the mine-water pools had become strongly acid in character and carried large amounts of sulphates in solution. This occurred in the following way. When the original sulphide water was removed, shafts sunk, and mining carried on over a number of years, ferrous sulphate was formed by the spontaneous oxidation of the pyrite beds in contact with air admitted through churn-drill holes and mine openillgs. These products of oxidation can be seen today in the old mines of the district as crusts of crystalline sulphates

Jan 1, 1949

By R. S. Busk

TWO groups of binary alloys were prepared. The first group consisted of those elements relatively soluble in magnesium: Li, Al, Zn, Ga, Ag, Cd, In, Sn, Hg, T1, Pb, and Bi. These are predominately Group B elements. The second group consisted of most of the remaining metallic elements, all of which are relatively insoluble in magnesium: As, Au, Ba, Ca, Ce, Cu, Ir, La, Mn, Ni, Pd, Pt, Rh, Sb, Si, Te, Ti, W, and Zr. These are all Group A, transition, or rare earth elements. Alloys of the first group were prepared by melting sublimed magnesium and elements as pure as obtainable in a graphite crucible, using a standard flux1 for protection. The melts were poured into a small graphite mold to produce 3xY4 in. diam slugs. Spectroscopic analyses were made of each slug. In all cases, the total impurity content was less than 0.05 pct. The slugs were then heat treated 5°C below the eutectic temperature until complete solution, as judged metallographically, was obtained. Filings from the center of the slugs were compacted at room temperature into 1/2 in. long x 1/4 in. diam cylinders, sealed in pyrex under 1/2 atm of purified argon, annealed at the heat-treating temperature for 1 hr, and quenched in water. After the X-ray diffraction patterns were made, each compact was analyzed chemically for the solute. It was found that the Mg-Li alloys reacted during filing to reduce the Li in solid solution. For this series of alloys, the slug was cold worked, annealed in purified argon, the surface dissolved in dilute HC1 to a depth of about 0.005 in., and the X-ray exposure made on the fresh surface. This procedure resulted in sharp, continuous lines and in true values for the lattice parameters. The alloys in the second group were prepared in the same way except that about 1 wt pct of each solute was added. Since the solubility limit was exceeded, not all the solute dissolved. These alloys were analyzed spectroscopically only to confirm the presence of the solute. The diffraction patterns were of the back-reflec-tion type using the characteristic copper radiation* and a flat cassette. Both the film and the specimen were rotated. The specimen temperature was maintained at 25" ± 1°C. The films had sharp lines; their diameters were measured directly to 0.01 cm. The parameters were calculated from the line diameters using a modified version of Cohen's extrapolation method.2 his method effectively eliminates systematic errors such as film shrinkage, inaccurate alignment, and non-precise film-to-speci-men distance. The random errors were minimized

Jan 1, 1951

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

By A. Kenneth Schellinger, O. Cutler Shepard

Overgrinding, Or the breaking of ore particles into sizes smaller than required for liberation, is a first-magnitude problem in grinding for concentration processes. The conventional ball mill-classifier circuit used in commercial grinding produces mineral particles that are all ground to pass an upper size limit, while the proportion of very fine sizes, or slimes, is uncon rolled. Calculations based on measurements made in commercial plants which were grinding ore for flotation concentration indicate that about go pct of the energy used in grinding is consumed in useless overgrinding.1 In addition to accounting for the consumption of a large amount of energy, over-grinding is detrimental to the flotation process. Overground slime coats larger partitles and inhibits selective action by the flotation reagents on particles which by themselves could be separated very well. A further difficulty is that extremely finely ground minerals do not respond satisfactorily to flotation-collecting reagents, and both selection and recovery decrease markedly with particles finer than 5 to 10 microns.2 conventional classifier-ball mill circuits inherently tend to overgrind the valuable minerals more than the gangue minerals because most valuable minerals are of higher density and are softer and more friable than the gangue minerals. The rem- edy to this situation would seem to be a separating mechanism which would separate free valuable mineral particles regardless of size and density considerations and a mechanism which would operate in the grinding mill so that valuable mineral grains would be removed from the grinding machine as soon as they are freed. A consideration of these ideas led the authors to the conception of a simultaneous grinding and flotation mechanism. So far as the authors are aware, such a machine has never been tried, even on a laboratory scale. Development oF Apparatus and Experimental Method Apparatus development went through several stages before a really workable machine was devised. Work was started with a large cast-iron mortar 7½ in. in diameter and 8 in- deep. A cast-iron disk Was mounted on the end of a vertical shaft SO that it fitted loosely into the bottom of the mortar. On top of the disk four radial cleats ¼ in. high were bolted to agitate the ball charge as the disk, or impeller, rotated. The vertical shaft was driven at 100 rpm by a mechanism improvised from an old coffee-mill type of grinder. A hole was drilled into the bottom of the mortar through which compressed air was introduced for frothing. In operation, more charges of Ore which had been crushed to pass an 8-mesh screen were placed in the bowl along with water and suitable reagents. The impeller was then started and the compressed air turned On so that the ore was ground while being exposed to air bubbles. Mineralized froth was

Jan 1, 1949

By H. J. Modi, P. L. De Bruyn

IN recent years interest has been aroused in flotation rate studies both from a technical and a more practical aspect. With increasing fineness of grind becoming a necessity in treatment of low grade ores, difference:; in flotation rates of various size ranges will assume importance in the evaluation of overall recovery results. Theoretical analysis of flotation kinetics was initiated by Schuhmann,' who sought to define flotation rate of a given mineral and specified size range in terms of certain physical variables of the flotation system. Using as a basis for his analysis the bubble-particle encounter hypothesis previously advanced by Gaudin,' Schuhmann writes for the rate of flotation (R,.) of particles of average size x, Rx = PcPa(x) V F. [1] In this relation Pc and Pa represent, respectively, the probability of collision between particle and gas bubble in a given time interval and the probability of adhesion of particle to the bubble after collision. Rigorously speaking, the term PC does not have the significance of a probability factor, since it depends on the time interval and therefore can be numerically greater than unity. The term c(x) expresses the concentration of particles of an average size x in the flotation cell in weight per unit volume of pulp, and V denotes the total volume of pulp in the cell. The product, P,c(x)V, defines the frequency of particle-bubble collisl.on, and when it is multiplied by P,, the number of collisions per unit time resulting in successful adherence is obtained. The factor, F, measures the froth stability and is introduced because after fruitful bubble-particle contact some particles may still riot reach the froth launder. Coalescence of bubbles in the froth column or too short a bubble life could account for this effect. Schuhmann also introduced the concept of "specific flotation rate" (Q) to allow for the relative abundance of different mineral species in the flotation pulp. The specific flotation rate, having the dimension of reciprocal time, is defined as Eq. 2: Qx =Rx/c(x)v [2] and it also follows from Eq. 1 that Qx = PcPaF. [31 The theoretical aspects of flotation kinetics were also considered by Sutherland,3 who tried to improve Eq. 1 by expressing the probability factors in terms of such physical variables as particle radius (x), bubble radius (X), velocity of bubble relative to particle (v), the number of bubbles per unit volume of pulp (N) and the induction period (A) necessary for air-mineral adhesion. According to Sutherland's treatment, the rate equation, Eq. 1, transforms to R, = [3pX x v N] [sech2 (3 v ?/4 X) ] ?c(x) V [4] where the first term in brackets on the right-hand

Jan 1, 1957

By D. B. Carson, D. L. Katz

Natural gases under pressure form crystalline hydrates with water. Experimental data are reported on four-phase equilibrium for the methane-propane-water, methanc-pentane-water, and methane-hexane-water systems. Temperatures and pressures for equilibrium between gas, water-rich liquid, hydrocarbon-rich liquid, and hydrate were measured, as well as the percentages of methane and propane in the hydrate. The data indicate that natural gas hydrate behaves as a solid solution and that pentanes and heavier hydrocarbons do not enter into the solid phase. Vapor-solid equilibrium constants are presented that permit the approximation of the conditions for hydrate formation, from the composition of a gas. Introduction Natural gas hydrates have been the object of considerable research in recent years, because of the trouble they have caused in the natural-gas and natural-gasoline industries. Natural gas hydrates are white crystalline compounds of water and gas, which, under pressure, exist at temperatures considerably above the freezing point of water. Because of the relatively high temperatures at which the hydrates exist, they become a nuisance in high-pressure gas operations where water is present, since their formation causes partial or complete plugging of valves and pipes. From a practical standpoint, the trouble incident to hydrate formation has been solved by dehydration of the gas before it enters the plant or pipe line, or by other remedial measures. Extensive work on gas hydrates in the latter part of the nineteenth century, by Villardl and others, gave data on methane hydrate and on ethane hydrate. Schroeder summarized these theoretical studies2 Hammerschmidt3 introduced the information on gas hydrates to the gas industry and Dcaton and Frost4-6 gathered a considerable number of data on the behavior of natural gas hydrates. Recent papers7,8 have extended both the data and the theory of these hydrates. At present, data are available on tem-peraturcs and pressures of hydrate forma. tion for pure methane, ethane, propane, n-butane, methane-ethane mixtures, meth-ane-propane mixtures, methane-butane mixtures, and some Is natural gases, all- in the Presence of excess water. However, up to the Present time no method has been available for predicting the temperature at given pressures or the pressure at given at at which natural gases will form hydrates if hydrates with water. 'This Paper Presents data on the methane-propane-water, methane-pentane-water and methane-hex-""e water systems in four phases, one of which is the hydrate. he application of the phase rule to the methane-~ro~ane-water data Proves that hydrates behave as solid solutions. The data of this paper and those in the literature were used to develop vapor-solid. equilibrium constants,8 which permit prediction of conditions of hydrate formation from the analysis of a natural gas.

Jan 1, 1942

The great Bingham Canyon mine and mill complex of Magna, Arthur, and Bonneville had their beginning in a 300 ton per day operation which was commenced by Daniel C. Jackling and his associates in 1904 as the Utah Copper Company. This small plant was quickly expanded to 500 tons per day and eventually Kennecott, the Utah Copper successor, was processing about 45,000 tons per day of ore in each of its mills, Magna and Arthur. Finally, in 1966, the Bonneville crushing and grinding plant was added to produce a slurry of ground ore for flotation concentration at the expanded Magna and Arthur mills. Bonneville is one of the few railroad dump primary gyratory crushing plants constructed in recent years; another which is described in this Monograph is Cuajone in Peru, started in 1976. Two others are the Climax Henderson unit and the Phelps Dodge Tyrone crusher, which has been fed by trucks since it commenced operation. The Bonneville, Henderson, and Cuajone crushers are all located considerable distances from their mines, hence the use of rail haulage. Shovels in the Bingham pit are 8 to 15 yd capacity; and ore is sent to the Bonneville crusher in 100 ton rail cars, where they are dumped, two at a time, by a hydraulically driven rotary dumper. Ahead of the 54 in. by 74 in. Allis-Chalmers gyratory are two grizzlies, each 27 ft 9 in. (8.46 m) long with bar spacings of 4 ½ in. (114 mm) at the feed end and 6 in. (152 mm) at the discharge end. Grizzlies are sloped at 38. Oversize rocks are re- moved from the crusher pocket by an overhead crane, and the crusher produces a nominal 6 in. (152 mm) product. Rail cars are dumped in the direction away from the crusher, and water sprays at the dumper level and in the rock boxes are used for dust suppression. There are two 72 in. by 19 ft (1.83 m by 5. 8 m) apron feeders under the 500 st cr'ushed ore pocket and two more under the grizzly undersize pocket. Feeder speeds are variable up to 72 fpm. Grizzly undersize is dropped to the feeder pocket by three shelves forming a rock ladder attached to the pocket walls. Each pair of apron feeders, one for crushed ore and one for grizzly undersize, discharge to a 72 in. (1. 83 m) picking belt for tramp removal; this belt feeds one of two 54 in. (1.37 m) feeder conveyors which in turn discharge to a 60 in. (1. 52 m) takeaway belt operating at 175-300 fpm (53.3 - 97.4 mpm). Both picking belts are equipped with metal detectors and tramp iron magnets, and a separate 30 in. (760 mm) conveyor is used to remove tramp material.

Jan 1, 1978

By G. L. Hassler, E. Bruner

An apparatus and method is described whereby the relation between saturation and capillary pressure under decreasing saturation can be quickly determined for small core samples. The initially saturated core is cen-trifuged at increasing rates and the average saturation is measured at each rate with the aid of a stroboscope device. A theory and calculation procedure is given whereby the accelerations and saturation values can be converted into a true curve of capillary pressure versus saturation. The present work differs from previously described use of the centrifuge to obtain capillary-pressure curves of core samples3 in that the core is centrifuged alone, so that the whole range of saturations required by the properties of the sample and the radially varying centrifugal force occurs within the sample. The calculation procedure nevertheless secures correct results from simply obtained values of the average saturation. The apparatus described was developed in connection with studies of capillary pressure and wetting of oil-field rocks, and the paper provides the apparatus background for a previous publication' dealing more fully with the uses and implications of the data in petroleum engineering. The method is applicable, however, to general combinations of immiscible fluids and porous or comminuted solids, and may be useful in other fields. Introduction The importance of capillary phenomena in determining the behavior of liquids in porous media has long been recognized. The concept of capillary Pressure was early formulated, and in recent Years discussions of the application of the laws of capillarity to oil field problems have appeared.l-3 One reason why progress in the application of capillary phenomena to the solution of problems of reservoir mechanics has not been more rapid has been the lack of a, satisfactory technique for measuring the capillary pressures in reservoir rocks. It is the principal purpose of this paper to describe a method by which the relation between capillary pressure and saturation can be determined for small consolidated core samples which have been extracted and resaturated, Capillary Diaphragm Perhaps the most direct way to measure the capillary pressure in a rock is to make piezometric contact with the wetting liquid through a diaphragm of porous material with high displacement pressure saturated with the liquid in question.3-6 The liquid in the diaphragm assumes the pressure of the liquid in the core, which can then be measured with a manometer, while the atmospheric air is prevented by capillarity from entering the diaphragm and affecting the pressure measurements as long as the displacement pressure of the diaphragm is not exceeded. With an apparatus of the type illustrated schematically in Fig. I, the relation between capillary pressure and saturation can be measured for a small core sample by starting with the core saturated 100 per cent, applyillg successively greater suctions

Jan 1, 1945