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By N. A. Gjostein

The real and apparent dependence of the surface self-diffusion coefficient, Ds, of gold on crystallo-graphic orientation has been investigated by isolated scratch smoothing and grain boundary grooving experiments. No systematic variation of Ds with orientation was found along the [100] zone. In view of the experimental scatter in these results, however, only aniso-tropy ratios, p, greater than -1.3 would be detectable. Anomalously low values of Ds were found on the [0ll] and [110] zones when the diffusion direction lies along (the scratch or grain boundary groove is perpendicular to) these zones. By comparing these results with a previous y plot study of gold, it is hypothesized that this is not an intrinsic diffusivity effect, but rather it results from a decreased driving force for mass transfer at orientations where the y-plot curvature is negative and appreciable in magnitude relative to y. The characteristics of this type of anisotropy do not appear sEVERAL investigations1-7 have dealt with the anisotropy of the surface self-diffusion coefficient, Ds, of metals as measured by the decay of single or multiple to be in agreement with the hypothesis of fast diffusion along ledges in the surface structure. In certain cases the Fourier method of determining Ds, described in Paper I of this series9 was compared with the second moment analysis in order to determine whether or not the method of analysis affects the anisotropy ratio. While the type of analysis does influence the magnitude of the anisotropy ratio, this effect is not associated with the shape of the profile, but rather is connected with the iterative volume diffusion correction used in conjunction with the second moment analysis. It is further demonstrated that the volume diffusion correction can be too large by either method when the y-plot curvature effect is large. Certain orientations exhibit an unusual temperature dependence of Ds ; this is discussed in terms of the y-plot curvature effect and/or the influence of impurities on Ds . scratches in the surface. Depending on the system investigated, both large1"3'6'7 and small4'5 anisotropies in Ds have been found. Although adsorbed impurities undoubtedly play an important role in determining the magnitude of the anisotropy,8 there are other factors that can give rise to an apparent anisotropy in Ds, thus obscuring the true nature of an intrinsic anisotropy in

Jan 1, 1968

By FRANK RIEBERS

IN DISCUSSING the selection of a geophysical method, much of what the writer will say is applicable to any of the various methods and to their use in prospecting, whether for oil or for other minerals. Specifically, the writer's subject is the problem, particularly pressing on the Pacific Coast, of checking the rising cost of the discovery of new fields by the proper employment of geophysical methods. Table I was prepared from statistics published by the California State Milling Bureau. The writer selected five of the largest a:ld most active companies operating in California, segregated the number of wild-

Jan 1, 1930

By A Allen, J. A. Garcia

The large increase in the wages of mine workers makes it imperative that all factors tending to limit production per miner be eliminated, if possible. The trolley and storage-battery locomotive, mining machine, hoisting engine, mining system, etc., have increased the possible output per man. But this output is still limited by the capacity of the hoisting engine and cable and by the size of the mine car. The superiority of skip hoisting in metal mining is shown by its almost universal adoption. By varying the size of skip and the rope speed, any desired tonnage can be secured and, since frequently only one kind of material is handled and breakage is unimportant, the loading of the skips can be easily and cheaply effected from bins into which the cars are dumped. The cars may then be designed to fit the conditions in the mine instead of being a compromise between hoisting and mining conditions, very likely suiting neither. In coal mines, the usual practice has been to hoist the car to the surface, either on platform or self-dumping cages. Of late years the number of skip-hoisting plants in coal mines has been rapidly increasing, but there is more or less inertia to overcome in establishing so radical a change in practice; also the earlier skip operations were not uniformly successful. As a result, it became evident that metal-mining practice would require radical modification before skip hoisting could be successfully used in coal mining. Objections to Skip Hoisting Breakage Except where coal is used for coking, or for other purposes whcre breakage is unimportant, breakage of the coal is vitally important all the way from the face to the railroad car. The average selling price is seriously reduced when the percentage of screenings is increased. Following metal-mining practice exactly, the coal would be dumped into a deep pocket, cut off into a skip load by a measuring hopper or other means, dumped into a deep, narrow skip, and dropped into a bin at the top, all of which would contribute to excessive breakage.

Jan 1, 1922

By B. M. Larsen, T. E. Brower, J. W. Bain

A mass of circumstantial evidence is presented to indicate that the main source of alloy losses in open-hearth tapping is oxidation by air, with the steel apparently reacting with an amount of oxygen equivalent to about 30 times its own volume of air. The effect is erratic from heat to heat, depending largely on turbulence and distance of free fall of the stream of liquid metal. THIS paper reports another phase in a general study of oxidation in the open-hearth, various aspects of which have already been discussed in previous papers.1,2,3 It is based on experiments conducted on commercial, basic open-hearth furnaces at various times when opportunity offered, beginning about 1938, in an attempt to determine the mechanism and measure the amount of air oxidation during the flow of the tapping stream through the air into a ladle. This effect is closely connected, not only with alloy losses and control of chemical analysis, but possibly with various questions regarding inclusion content and steel quality and with a true picture of the deoxidation of steel in general. Thus it is believed that discussion of a large part of the results to date may be of assistance to others who are studying problems related to deoxidation or to steel cleanliness and quality. The literature reveals that strangely little attention has been devoted specifically to the effect of pouring steel through open air. Bardenheuer and Henke4 in 1939, in a paper devoted largely to overall losses of manganese in the basic open-hearth, reported that for a number of heats tapped into a tilted ladle held just beneath the runner so that the maximum distance of drop of the stream through the air was held to less than about 30 in., there was in most cases almost no loss of manganese from furnace to ladle. Hultgren,5 in 1945, showed that in high-carbon basic electric steel deoxidized in the furnace and containing no large inclusions as it left the furnace, large silicate inclusions, often rich in manganese oxide, were present in samples dipped from the pool in the ladle; also that the rise in content of large inclusions (above 0.01 mm dim) was greater when the metal stream was made to spray or flow in a turbulent manner. Our interest in this phase of oxidation in the open-hearth was first aroused by a study of alloy efficiencies on ladle additions. The frequency curves of manganese efficiency (net recovery of metal added in ladle) given in fig. 1 are of interest in this connection. Lowest, and also most variable (47 to 95 pct), recovery was in low-carbon heats having no coal additions to the ladle, progressing upward to a maximum average and least variable (82 to 97 pct) recovery in high-carbon heats killed with aluminum and silicon. Both coal additions and higher carbon contents in the tap stream, as well as additions of silicon and aluminum, help to lower the amount of manganese loss. But the most interesting point is the amount of manganese oxidized in many heats tapped at 0.50 to 1.06 pct carbon. It has been shown in a previous paper3 that the amount of oxygen in the tap stream in such heats is very small, usually below 0.01 pct, too small to oxidize any appreciable amount of manganese. Also, the fact that manganese should be oxidized at all in a metal solution containing such more stable oxide-forming elements as aluminum and silicon is an indication of some rather intense oxidizing effect such as could be caused by reaction with the air during ladle filling. Oxygen Balances from Tap Stream to Teeming Stream: Data in the form of alloy efficiencies, such as those shown in fig. 1, are not a proper measure of such a general oxidation effect. Having available

Jan 1, 1951

By I. A. Ettlinger

DEPRECIATION allowances have become firmly rooted in our income tax structure both by legislation and by court decisions. Secretary of the Treasury Morgenthau has recently stated before the Ways and Means Committee of the House: A taxpayer who did not provide for depreciation of assets used in his business could not long remain in business and the income tax should be levied on the assumption that he will provide for such depreciation. The 1934 Revenue Act has now been enacted, the depreciation provisions of the 1932 Revenue Act having been carried forward without change. However, there has arisen, in connection with the preparation of this new Revenue Act, a determination by the Treasury Department to make certain changes in the future administration of the depreciation allowances that will affect all corporations, and mining companies in particular.

Jan 1, 1934

By George Watkin Evans

The need for mechanization in coal mining wherever it can be successfully introduced is evident when we take into account the fact that during recent years the coal-mine employee's wages have increased and his hours have been shortened. Coal is no longer King, and a brief study of statistics covering the production of coal, petroleum and natural gas and the development of water power will convince anyone that coal has had very powerful competitors during the past 10 years. Petroleum production in the United States between 1918 and 1936 increased from 100 per cent to 309 per cent. Water power increased from 100 per cent in 1918 to 270 per cent in 1936. Bituminous coal slipped from 100 per cent in 1918 to 75 per cent in 1936. The anthracite industry slipped from 100 per cent in 1918 to 50 per cent in 1932, but climbed to 55 per cent in 1936. In 1918, the State of Washington produced a little over four million tons of coal and the production of California oil, at that time, was a little less than one hundred million barrels per year. By 1929 the production of California oil had increased to a little over 292,000,000 bbl. or nearly 300 per cent as compared with 1918. By 1934 the production of coal in the state of Washington had dropped to about 1 1/3 millions of tons per year. Our neighbors in British Columbia had their coal production reduced from a little over 2 1/2 million tons in 1920 to less than 1 1/3 million in 1933. The North western Improvement Co., a subsidiary of the Northern Pacific Railroad Co., owns coal properties in various parts of Washington and Montana. The Northern Pacific is one of the very few railroads using any considerable quantity of coal in its operations. Practically all of the American railroads in the Pacific Northwest, except the Northern Pacific, are using cheap California fuel oil or electricity. This problem, of course, is not a new one in coal mining, and all of the progressive coal companies in the country have tried to meet competition from an overproduction of petroleum and an overproduction of natural gas as well as enlarged developments of hydroelectric plants.

Jan 1, 1938

By George Watkin Evans

The need for mechanization in coal mining wherever it can be successfully introduced is evident when we take into account the fact that during recent years the coal-mine employee's wages have increased and his hours have been shortened. Coal is no longer King, and a brief study of statistics covering the production of coal, petroleum and natural gas and the development of water power will convince anyone that coal has had very powerful competitors during the past 10 years. Petroleum production in the United States between 1918 and 1936 increased from 100 per cent to 309 per cent. Water power increased from 100 per cent in 1918 to 270 per cent in 1936. Bituminous coal slipped from 100 per cent in 1918 to 75 per cent in 1936. The anthracite industry slipped from 100 per cent in 1918 to 50 per cent in 1932, but climbed to 55 per cent in 1936. In 1918, the State of Washington produced a little over four million tons of coal and the production of California oil, at that time, was a little less than one hundred million barrels per year. By 1929 the production of California oil had increased to a little over 292,000,000 bbl. or nearly 300 per cent as compared with 1918. By 1934 the production of coal in the state of Washington had dropped to about 1 1/3 millions of tons per year. Our neighbors in British Columbia had their coal production reduced from a little over 2 1/2 million tons in 1920 to less than 1 1/3 million in 1933. The North western Improvement Co., a subsidiary of the Northern Pacific Railroad Co., owns coal properties in various parts of Washington and Montana. The Northern Pacific is one of the very few railroads using any considerable quantity of coal in its operations. Practically all of the American railroads in the Pacific Northwest, except the Northern Pacific, are using cheap California fuel oil or electricity. This problem, of course, is not a new one in coal mining, and all of the progressive coal companies in the country have tried to meet competition from an overproduction of petroleum and an overproduction of natural gas as well as enlarged developments of hydroelectric plants.

Jan 1, 1938

By J. M. W. Mackenzie, C. J. Wood

In attempting to describe the size classification performance of a hydrocyclone, most workers have elected to use either an equilibrium orbit theory or an non-equilibrium orbit theory. The equilibrium orbit theory has been used by the majority of workers including Lilge,' Bradley; and Yoshioka and Hotta. In applying this theory, it is argued that particles in the body of a hydrocyclone attain an equilibrium radial position where the drag force on the particle resulting from the inward radial fluid velocity is balanced by the outward centrifugal force caused by the tangential component of fluid flow. When considered over the full height of the hydrocy-clone, attainment of this radial equilibrium orbit results in the particle following a conical equilibrium envelope. It is then argued that if this envelope lies outside the envelope of "zero vertical velocity," the particle will report to the underflow, while if the equilibrium envelope lies inside the envelope of "zero vertical velocity," the particle will report to the overflow or vortex finder product. The d50-sized particle which reports in equal quantities to the underflow and overflow is assumed to correspond to particles whose equilibrium envelope is coincident with the envelope of "zero vertical velocity." In considering the equilibrium orbit theory, it is apparent that the horizontal position of the particles in the feed inlet pipe should have no effect on their ultimate destination on the hydrocyclone. Each particle should attain an equilibrium position which depends on the density, size, and shape of the particle; the density and viscosity of the fluid; and the flow patterns within the hydrocy-clone. The nonequilibrium orbit or unsteady state theory has been largely developed by Rietema4 and Mizrahi.6 Mizrahi has listed four main objections to the equilibrium orbit theory. These objections center on the short residence time in the hydrocyclone, the fact that the experimental classification curve is much less sharp than is theoretically predicted, and the absence of negative efficiency conditions in hydro cyclones operating on a feed material which is much finer in size than d50. Proponents of the nonequilibrium orbit theory argue that for a particle to discharge with the underflow it must have sufficient outward radial velocity to reach the downward-flowing region close to the hydrocyclone wall in which the flow lines are parallel to the wall and the ratio of vertical to radial velocity is constant. It is then postulated that a d50 particle entering the cyclone at the center of the feed inlet will just reach this downward-flowing region as it reaches the apex. Thus for uniform distribution of particles across the feed inlet, half the d50 particles—that is, those injected in the half of the inlet area nearest the cyclone wall —will report to the underflow while those injected in the other half will not reach the downward-flowing region and will be carried inward to the center of the cyclone and thus report in the overflow. The exact thickness of the down-ward-flowing region of fluid adjacent to the outer wall of the hydrocyclone is uncertain but Mizrahi considers it to be equal to the apex radius minus the air core radius. Particles larger than d50 have a greater outward centrifugal force acting on them than the d50 particles and may reach the wall even if injected at a distance from the wall greater than Di/2 (Di is inlet diameter). Conversely, particles smaller than d50 may not reach the wall even if injected at a distance less than Di/2 from the cyclone wall. Since the equations put forward by the proponents of both theories yield approximately the same values of d50, it is not possible to decide between these theories by measurement of d50. It should be possible however to examine the theories by injecting a small stream of solids into the feed inlet of a hydrocyclone running on clear water. If the efficiency or classification curve is measured for various horizontal injection positions, then the curves should be coincident if the equilibrium orbit theory holds. If, however, the unsteady state theory describes the cyclone operation, then the classification curves should show finer d50 sizes for particles injected close to the cyclone wall. Experimental A 6-in.-diam hydrocyclone with geometry as in Figs. 1 and 2 was used. Quartz particles were injected as a 50% by wt pulp via an 1/8-in. steel probe. For each in-

Jan 1, 1971

By C. J. Bonnecarrere, P. A. Jr. Bloomer, J. Hunter

Since 1941 not more than 15 per cent of aii wildcat wells drilled in Louisiana have been successful. This figure is not too discouraging, especially in view of the fact that during the same period approximately 80 per cent of all wells (average 701 wells per year) drilled in or adjoining producing areas were successfully completed. However, a somewhat discouraging aspect concerning Louisiana's wildcat completions is that in 1942 only 12 out of 20 fields found were oil fields; in 1943, only 13 out of 24; and in 1944, only 8 out of 21. This means that an operator has less than a 50:50 chance of finding an oil field. In other words, it is more likely that he will find a gas-condensate field. This fact has become increasingly evident as the data for each year are compiled and compared. Yet, strangely enough, operators are reluctant to accept this fact and plan accordingly. Certainly they cannot continue to search for oil fields and, for the most part, end with shut-in gas-condensate wells. The reason for shutting in is that there is no immediate and satisfactory market for all the gas being discovered. The solution, and it is a very satisfactory one, as experience in Louisiana has already demonstrated, is cycling. The advantages of this type of operation of gas-condensate fields, especially where the entire field is pooled and unitized, should be self-evident. These advantages will become increasingly more important, especially now that legislatures, regulatory bodies, royalty owners, and others are beginning to realize that "reasonable development" in a field is not indicated by the number of wells completed, but rather by the method of development. To complete a great many wells that can produce at only, partial capacity is waste. Producing hydrocarbons from these wells and not subjecting them to processing by surface operations, in order to recover all liquefiable hydrocarbons, is waste. Permitting bottom-hole pressures to drop or gas-oil ratios to increase because of excessive production is waste. Loss of leases and of wells because of inability to produce or to meet lease obligations is waste. Therefore, it is becoming increasingly apparent that a reservoir entirely pooled and unitized with a minimum number of completions, but over which there is exerted careful control so as to provide maximum recovery and the processing of all hydrocarbons brought to the surface, is the only logical and reasonable method of developing a gas-condensate field. Oil Production Twenty-one new fields, including three rediscoveries, were found during 1944. Three of these new discoveries—Delhi in northeast Louisiana, Good Hope in southeast Louisiana, and West Tepetate in southwest Louisiana—should add considerably to Louisiana's oil reserves. The other discoveries were either gas condensate or of apparently little consequence. State-wide productioll for 1944 was 1366 million barrels of crude oil and condensate; amounting to about 8.1 per cent of the

Jan 1, 1945

By Gerald C. Mathis

FERGUS County, Mont., shown in Fig. 1, is known for its once famous gold mines near the old towns of Gilt Edge, Maiden, and Kendall. But at Heath, a small farming community near the foot-hills of the Snowy Mountains, thousands of tons of gypsum have been mined since 1916. The value of the gold taken from the mines of Fergus County runs into millions of dollars; the value of the products made from the gypsum of the Heath mine is also in the millions. The gypsum mine will continue to produce for many years, while the gold mines are closed today.

Jan 2, 1953

By AIME AIME

PRACTICALLY all the Section delegates as well as a sprinkling of Institute officers and mere members were on hand for the annual business meeting of the Institute on Monday afternoon of the Annual Meeting. President McAuliffe presided, with C. H. Mathewson as associate chairman. Americo Albala was first called on to discuss the work of the Pan American Institute of Mining Engineering and Geology. C. Will Wright said that the need for such an organization was evident, and a resolution was passed asking the incoming President to appoint a committee to investigate cooperation with that body, of which Mr. Albala is executive secretary.

Jan 1, 1943

By R. R. Vandervoort

The creep behavior of Nb(Cb), Ta, Mo, and W was determined under conditions of constant atomic dif-fzisivity, constant stress to elastic modulus ratio, and nearly equivalent grain size, and the steady-state creep rates obtained from these tests were correlated with calculated stacking fault energies for the metals. These results, in conjunction with similar data for several fccMetals,13 suggest that stacking fault energy may influence the creep strength ofbcc metals. The interrelationship between steady-state creep rate, subgrain size, and stacking fault energy was examined. It was found that the subgrain size for a given creep stress, increased as stacking fault energy increased, but that this relationship did not cormpletely account for the effect of stacking fault energy on creep rate. The crystallography and energetics of stacking fault formation in bcc metals has been discussed by a num-ber of authors,1-5 and impurity stabilized stacking faults on (112) planes have been observed in Nb,6,7 w,8,9 Fe,] and V" by transmission electron microscopy. However, a crucial question is whether or not stack-ing faults influence the mechanical strength of bcc metals. Potentially, stacking faults could increase strength by reducing the mobility of the partial dis-locations bounding the fault, by acting as barriers to slip dislocations, and by retarding the climb of dislo-cations during high-temperature deformation. The objective of this study was to seek a correlation be-tween creep strength and stacking fault energy for several bcc metals; namely, Nb, Ta, Mo, and W. The creep behavior of most polycrystalline metals and alloys at high temperatures and moderate stresses can be described by the following relation:11,12 im=Af(s) where i, = minimum creep rate, A = constant, j(s) = a function involving metallurgical structure, a = applied stress, E = average elastic modulus at the test tempera-ture, w = constant (equal to 5 for most pure metals), D = diffusion coefficient. One factor in the structure function F(s) which sig- R. R. VANDERVOORT, Member AlME is Research Metallurgist, Process and Materials Development Division, Chemistry Department, Lawrence Radiation Laboratory, University of California, Livermore, Calif. Manuscript submitted February 28, 1969. IMD nificantly affects the creep resistance of fcc metals is stacking fault energy, and creep rate has been shown to vary directly with stacking fault energy to the 3.5 power." In the latter investigation, four fcc metals of widely different stacking fault energies (Ag, Cu, Ni, and Al) were creep tested at a constant stress to modulus ratio of 1.21 x 10-4, at a constant diffusivity of 2.7 x 10-12 sq cm per sec, and at nearly equivalent grain sizes of about 0.7 mm. The creep data were then correlated with stacking fault energies. In the present study, a similar procedure was followed. All materials used in this work were consolidated by powder metallurgy techniques. Impurity contents in the as-received materials are listed in Table I. Chemical analyses showed that no measurable contamination of the test specimens occurred during pretest annealing treatments or creep testing. Specimens with a gage section 0.75 by 0.125 by 0.050 in. were creep tested in tension in a vacuum of less than 10-9 torr. Deformation at temperature was measured by tracking fiducial marks on the gage section of the specimen with an optical comparator. Optical deformation measurements also permitted observation of the macroscopic characteristics of the deformation Table I. Typical Specimen Impurity Content, ppm Nb Ta Mo W C 45 10 155 6 O 185 30 4 10 N 30 6 3 2 H 5 I 1 <1 als 3 10 2 15 Ca <5 I3 5 Cr 5 <3 10 <5 Cu 10 50 2 15 Fc 10 10 150 35 Ni 2 150 20 <5 Si <I0 1 3 <10 Ta 100 Ti 10 8 1 Zi 15 50 1 3 Table II. Test Conditions for Constant Stress-Modulus Ratio of 6 X 10.&apos; and Constant Diffusivity of 2.7 X 10-12 sq cm per see, and Grain Size Values for the Given Pretest Annealing Treatments Literature references Pretest Annealing for E and D Treatment Stress, Temperature, ___"&apos;Values__ Grain Tempera-Metal psi "C E D Size, mm ture, .C Time hr Nb 745 1525 14 15 to 17 0.83 1650 I Ta 1220 1770 18 19.20 O.91 1800 I Mo 1975 1630 18 21 0.77 2200 I W 2140 2265 18 22 040 2400 5

Jan 1, 1970

By James W. Wade

THIS discussion of the future supply of lead refers only to the next ten-year period. Beyond that no prediction can be made that would be of sufficient accuracy to serve any purpose. When any commodity is by an unsatisfied demand forced to a higher price level, and the profits of those who are able to produce the commodity are greatly increased, there is awakened a new interest in the commodity on the part of both producer and consumer. The consumer attempts to reduce his costs by developing substitutes or assisting in the development of new sources of supply. Lead is now going through this stage. The lead producer is increasing development expenditure and searching for new mines. The consumer is developing substitutes as illustrated by the increasing use of zinc oxide as a pigment, and the use of dry batteries in radio sets. The search of the Ford interests for lead, by prospecting in Missouri and Idaho, is induced by the effort to develop a new source of supply. Considering the index figures of the Department of&apos; Commerce, it is evident that lead is on a substantially higher price level. For February, 1926, the index figure of lead is 209, while the general commodity price index is only 156, 100 being taken as the base for 1913.

Jan 1, 1926

By AIME AIME

AN outstanding session of the Annual February Meeting was one held under the joint auspices of several groups on Tuesday afternoon, Feb. 22, as a symposium on the question of preserving peace in the postwar era through the control of the supply of sundry minerals vital to the Axis nations. D. H. McLaughlin and R. C. Allen were associate chairmen. And C.K. Leith led the discussion.

Jan 1, 1944

By A. T. Aldred

HIS note reports the existence of several new scandium intermetallic compounds of the A2B and AB stoichiometries where the A element is scandium and the B element is from group VIII or IB of the periodic table. Alloys were arc melted from high-purity materials using 2-g charges. Typical lot analyses for the materials used are given in Table I. Chemical analyses were not made on the alloys as there was little or no weight loss on melting. After annealing the specimens for 14 days at 600°C and water quenching, metallographic and X-ray techniques were used to determine the occurrence of the various phases. Table II lists the relevant data for the A2B phases. The Ti2Ni structure,2 has ninety-six atoms in a face-centered cubic cell, whilst the Al2Cu structure is body-centered tetragonal.3 The occurrence and stability of the various A2B type phases in transition metal alloy systems has recently been considered by Nevitt.4 The Ti2 Ni-type phases occur over a relatively narrow range of radius ratios from 1.14-1.26 with A partners mainly from the titanium group and B partners predominantly from the cobalt group. This suggests that both a favorable relative size of the atoms and a favorable electron concentration are necessary conditions for the existence of this structure type. A12Cu-type phases involving Ti group elements as A partners are formed with B partners from the cobalt and nickel groups, predominantly the latter, and have a higher radius ratio (1.27-1.29) than the Ti2Ni-type phases. The occurrence of both Sc2Pd and Sc2Co can be rationalized in terms of these considerations but SczNi is anomalous because of its high radius ratio. The radius ratios are calculated from the Goldschmidt CN12 radii and it might be postulated that the scandium atom is reduced in size in the compound. Unfortunately the apparent atomic size of scandium in the compound cannot be calculated from a knowledge of the structure and lattice dimensions. The crystal structure of the Sc2Au compound could not be identified from the powder diffraction pattern, but it is interesting to note that it is isostructural with a series of A2Au phases where A is a rare earth element.5 Pertinent information concerning the AB phases found in this investigation is given in Table 111. All the phases have the ordered bcc CsCl structure. The occurrence and stability of CsC1-type phases have recently been discussed by Dwight,7 who concludes that the relative sizes of the atoms do not appear to be a controlling factor (the radius ratios of all the known phases vary between 0.985 and 1.439) but that the occurrence of the phase is predominantly dependent on the relative positions of the A and B partners in the periodic table. As a measure of the stability of a given CsCl phase it has been customary to define a lattice contraction on alloying, DAB — dAB, where DAB is the AB interatomic spacing as calculated from the Goldschmidt CN8 radii and dAB is the AB interatomic distance in the compound (= v3/2 ao). Values of the parameter DscB -dScB for the compounds found in this investigation are given in column 5 of Table III and are plotted in Fig. 1. If the lattice contraction does in fact reflect the stability of a phase then Fig. 1 shows two very interesting trends when the position of the B partner in the periodic table is considered. Firstly, within any given group the stability increases in order on going from the first to third long period. Secondly, within any given period the stability increases in order as the B partner is taken from the copper, nickel, and cobalt groups. In the case of the second long period, the lattice contraction reaches a maximum at ScRh and then decreases to ScRu. The same trend could

Jan 1, 1962

By Roy D. McLellan

Electrolytic production of cadmium at the Great Falls plant started in the first part of the year 1925. Prior to that time, an experimental unit had been in operation for a few months during the year 1922. At present the plant is the largest producer of cadmium, not only in the United States but in the world. Electrolytic cadmium is characterized by its very low content of impurities. The main impurities, which are lead, copper, zinc and iron, do not aggregate more than 0.05 per cent., leaving a cadmium content better than 99.95 per cent. The consumption of cadmium has increased rapidly since the war, with the growing use in the plating industry. Cadmium plating has come into prominence because of its connection with the automobile and radio industries. As a protective coating, a deposit of cadmium is superior to zinc. Cadmium is also used extensively in alloys. Important among these is the alloy with copper used in telephone and trolley wire. In proportion of 0.5 to 1.2 per cent., cadmium adds materially to the strength and wearing qualities. Practically all commercial zinc concentrates contain a small amount of cadmium. In the electrolytic zinc process, the cadmium that is dissolved in the regular leaching operation is removed from solution with metallic zinc, generally added in the form of zinc dust. This precipitate also contains the copper taken into solution and the excess of zinc dust, which is required in order to insure a complete removal of all impurities. The precipitate is treated by a separate process to recover the copper, cadmium and excess zinc. In order to make the metals more soluble and to render insoluble such impurities as iron, arsenic and antimony, the precipitate is roasted. The roasting is carried out in gas-fired McDougall furnaces, with a roasting temperature of about 700&apos; C. The roasted residue is leached with dilute sulfuric acid, spent electrolyte from the zinc electrolyzing, containing 10 to 12 per cent. free H2S04. The leaching is done in cylindrical tanks of 30 tons capacity, agitated

Jan 1, 1930

[Separates of all the Technical Publications published in 1928 are available at Institute headquarters. All the papers are on file in public, university and technical libraries, and when so indicated in the list, may be found either in the TRANSACTIONS or in the special volumes issued by the Institute] . METAL MINING TECHNICAL TRANS. PUBLI- TITLE AUTHOR MEETING VOLUME CATION AND PAGE* No. 50 .The Eötvös Torsion Balance D. C. Barton New York Method of Mapping Geologic Feb., 1928 Structure 52 A Cartographic Correction for the C. A. Heiland New York Eötvös Torsion Balance Feb., 1928 61. Liquid-oxygen Blasting at Chuqui- H. C. Schultz, New York camata, Chile F. K. Middleton Hunter Feb., 1928 76, 108, 123 64 Working Methods of Practical Geo- Hans Haalck New York physics Feb., 1928 65 Experiments with Eötvös Torsion P. W. George New York Balance in the Tri-State Zinc and Feb., 1928 Lead District 74 Gravity Surveying in Great Britain H. Shaw New York Feb., 1928 75 The Computation of Eötvös Gravity E. Lancaster-Jones New York Effects Feb., 1928 82 Earth-resistivity Measurements in W. 0. Hotchkiss, New York the Lake Superior Copper Country W. J. Rooney, Feb., 1928 James Fisher 83 Mining Methods and Records at the Roy W. Moore New York 76, 56, 90 United Eastern Mine Feb., 1928 98 Recent Results in Electrical Pros- Hans Lundberg New York

Jan 1, 1928

By Jim F. Lemons

INTRODUCTION A nickel doesn&apos;t buy much anymore. That&apos;s even true in the cost of recovering nickel -- the commodity. A 5[C] per pound (11 [c] per kilogram) increase in the nickel price won&apos;t cover a 10-percent increase in fuel costs for processing laterite ores, nor will it cover a 10- percent increase in wages for recovering nickel from a sulphide ore. It won&apos;t even cover a 5- percent increase in the capital cost necessary to mine and process ocean nodules. The basis for this statement and a quantitative analysis of the variations in costs to recover nickel from different types of nickel ores will be discussed in this paper. During the early part of the 1970&apos;s, the interest in nickel and its potential development was shifting from Canada to the Pacific Basin area. Canadian growth rate was slowing and each additional increase in nickel capacity was very expensive (Argall, G.O., 1970). Laterite ores were attractive to developers since they were easy to mine at a low labor cost, and technologies were being developed that would allow the recovery of nickel in marketable forms (Mohide, T.P., Warden, C.L., Mason, J.D., 1977). During the last decade, new laterite production was achieved in the Philippines and Botswana, and a significant increase in production from Indonesia and New Caledonia occurred. However, recent reports on nickel identify numerous problems plaguing the industry, i.e., high energy costs that are forcing shutdowns (INCO, 1980), (Mohide, T.P., Warden, C.L., Mason, J.D., 1977), high Canadian sulphide mining labor costs (Anon., 1976), and rapid inflation of nickel capital costs (Anon., 1981b). No definitive study has been made to determine how sensitive the total cost of production is to rising energy, labor, or capital costs, or to identify to what extent increased byproduct revenues could be used to offset these production costs. The purpose of this paper is to quantify the effects of increases in energy, labor, capital costs, and by-product credits on the future cost of production from nickel reserves and/or resources. METHODOLOGY OF ANALYSIS In order to evaluate the sensitivity of nickel production costs to changes in energy, labor, and capital costs, and in byproduct revenues, an analysis was made of existing and proposed nickel operations, their costs, and their sensitivity to change. The data for these analyses were obtained from numerous sources including international govern¬ment and university publications, professional journals, company reports, private communications, U.S. Bureau of Mines contracts, and estimates by Bureau of Mines personnel. These data were collected, in part, as an ongoing effort of the U.S. Bureau of Mines Minerals Availability System (MAS) to systematically measure and classify identified mineral resources according to their respective extraction technologies, economics, and commercial availability. The MAS Program is currently evaluating the present and potential availability of nickel to the United States in relation to mining, beneficiation, smelting, refining, leaching, transportation, infrastructure, environment, land use, labor, productivity, technology efficiencies, operating capacities, deposit life, and political factors. As a summary of this data, the following briefly describes nickel ore types as each relates to energy, labor, capital, and byproducts. Sulphide Operations Nickel is currently available from sulphide deposits. Roughly 55 percent of the current world nickel production in market economy countries is from sulphide ores, although only about 20 percent of the known land-based nickel reserves and/or resources are sulphides (INCO, 1981). In 1979, the Western World sulphide reserves were estimated at about 9.2 million metric tons of nickel (Little, A.D., 1979). Sulphide ores are typically mined underground, thus resulting in high mining costs; however, these ores can often be upgraded by flotation prior to final nickel recovery and require about one-third of the energy required to recover nickel from laterites (Anon., 1981c). In addition, nickel in sulphide ores may be associated with economically recoverable byproducts, which include copper, cobalt, gold, silver, and other precious metals. Laterite Operations Forty-five percent of current nickel production is from laterite deposits. Laterite deposits account for 80 percent of the known land-based nickel reserves and/or resources (INCO, 1981). In 1979, the Western World laterite reserves were estimated at over 35 million metric tons of nickel (Little, A.D., 1979). Nickeliferrous laterites are residual soils formed by weathering, typically under tropical or near tropical conditions. They can be easily mined by open-cut methods, but generally cannot be beneficiated; thus, recovery of nickel must entail processing of the total mined ore. Usually, these ores have high moisture content (20 to 30 percent by weight) that result in additional fuel costs to dry the ore. Nickel silicates, garnierites, are, processed by pyrometallurgical methods; limonite nickel oxides are normally leached. Byproduct cobalt is not recovered from most ferronickel operations, and is only recovered to a limited extent from many leach operations.

Jan 1, 1982

By W. L. Finlay

Size effects in quenching steel are particularly prominent and well recognized because of the existence of a critical cooling rate separating nuclea-tion and growth transformations, as exemplified by the formation of pearl-ite, from the shear type of transformation characterizing the martensite reaction. The absence of a similar, sharply demarcating cooling rate is characteristic of precipitation hardening systems and size effects in these are correspondingly less prominent. This paper is concerned with a size effect in high-purity, precipitation-hardenable alloys which appears not to have been recognized previously. This effect is believed to result from the thermal fluctuations which inevitably occur in quenching a specimen of finite size into a cooling liquid rather than from the existence of a critical cooling rate. QUENCHING "ANOMALIES" The precipitation hardening literature contains many references to so-called anomalous quenching and aging results. By "anomalous", investigators usually meant that their alloys did not consistently harden on aging after quenching according to some anticipated pnttern: not infrequently virtually no age hardening occurred with compositions known to be precipitation-hardenable. Perhaps the most prominent series of anomalies was reported over a 15-year period by Gayler and coworkers on aluminum-copper.l,2,3,4 Working with silver-rich copper alloy, Cohen5 secured some very interesting Rockwell F hardness changes of from 1-5 Rockwell points between the first peak and valley. Fink and Smith borrowed one of Cohen&apos;s specimens and, following his heat treatment and quenching practice to the letter, secured6 very different results in which only one peak was obtained and in which the hardness level at the aging lime of Cohen&apos;s first peak was 20 Rockwell F points higher. Fink and Van Horn7 encountered serious deviations in age-hardening high purity aluminum-zinc alloys. They stated that "the discrepancies were larger than might be expected from possible experimental errors in the Brinell readings." Geisler, Barrett and Mehl8 likewise encountered anomalous hardening results in aluminum-silver alloys, securing quite different age-hardening curves from two identical specimens. They stated that "the treatments of these samples had been the same, so that the different behaviors could be attributed only to accidental variations in the quenching practice."

Jan 1, 1950

By Dennis E. Stover

Grade control for in situ uranium leaching is maintaining, at desired levels, the uranium concentration in the pregnant lixiviant which feeds the extraction (ion exchange) circuit. This differs from grade control in conventional mining in that in situ grade control is imposed subsequent to rather than prior to leaching without the direct observation and control possible in conventional mining and milling. Grade control is a necessary engineering task through the entire project life, beginning with preliminary evaluation and design and ending only when the final wells are shut in. More specifically, this task is associated with designing the initial well field, maintaining productivity during well field operations, and planning well field expansions to compensate for the depletion of existing wells. Our aim during each of these phases is to maximize uranium productivity while ensuring that we are developing information necessary for future phases. Throughout this task, we remain cognizant of the fact that uranium productivity is a function of both the complex geologic and operational parameters of the project. A minimum list of significant geological factors must include the following: . Host formation mineralogy . Uranium mineralogy . Permeability . Ore quality (grade and thickness) Our concerns are not limited to average or general descriptions of these factors. The interactions among the factors as well as localized variations among them are far more important in obtaining high productivity from a deposit. For example, an ore body may display an average permeability of one darcy but have its high grade ore confined to regions with permeabilities less than 500 millidarcies. This relationship between ore grade and permeability must be understood early in the project life if the operation is to be successful. Similarly, the fundamental operating parameters for the well field include those items which enable us to distribute the lixiviant within the ore as well as the chemistry which will dissolve the uranium minerals. A minimum list of such factors will include: . Well design . Well efficiency . Well bore damage . Pattern size and shape . Oxidant type and concentration . Complexing agent concentration - pH The interactions between these variables and the complex geological variations create a situation which often defies quantitative analysis. Hence, we have adopted a statistical approach when correlating production history with these parameters and the quality of the correlations improves as the number of wells operated increases. However, the correlations are often specific to an ore body. We find that we must tailor the process to meet the unique geologic parameters of each ore body. PHASE 1: INITIAL WELL FIELD DESIGN Designing the initial well field is perhaps the most difficult phase because we have the smallest data base from which to predict performance. We will have conducted numerous laboratory tests and, perhaps, field pilot tests to select the lixiviant system and to characterize the geologic parameters. However, we will learn a great deal more about the deposit and its productivity during actual commercial scale operation. Fully acknowledging its limitations, we will, as a starting point, assume that uranium productivity will be directly proportional to the quality and quantity of recoverable reserves. Future deviations from this assumption will be plentiful and result from our limited understanding of the aforementioned interactions between and among the operational and geological factors. However, in planning the initial field, estimation of the quantity of leachable uranium present at levels above the economic grade and grade-thickness limitations is a priority task. (The concentration of uranium in the ore, the ore grade, is expressed in units of weight percent. The grade-thickness of the ore is the product of the average grade and ore thickness expressed in units of percent-feet.) These parameters provide a direct indication of the quality and quantity of ore. They are determined from gamma logging of closely spaced drill holes (50 to 100 foot centers being typical) accompanied by uranium core analysis or direct uranium assay logging. Gamma- logging measures gross gamma radiation which emanates from uranium decay products and indicates only the possible presence of uranium. Direct uranium assays are necessary to confirm this presence and the true location and quality of the reserve. From this information, an estimate of the in-place uranium reserve is formulated. Utilizing the above data, we begin to assess the quantity and quality of the leachable reserve. Analysis of self-potential and resistivity logs in conjunction with uranium logs will identify those zones which contain uranium and display apparent high permeabilities. Similarly, zones with extensive calcite cementation or dense clay sedimentation are located and uranium within such zones is excluded. (Interpretation of these logs should be confirmed

Jan 1, 1980