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By Irving Cadoff, Jack Medoff

THE linear thermal expansions of oriented single crystals of zinc were measured in the range from 20" to 416°C using a Leitz HTV optical lever differential dilatometer. The single crystals, supplied by Dr. J. J. Gilman, General Electric Research Laboratory, Schenectady, N.Y., were prepared from 99.999 pct-purity zinc, and were in the form of rods approximately 2 in. long by 0.15 in. in diameter having the following orientations: x = 1.5, 12, 20, 30, 37.5, 49, 60, 73, and 82 deg where x is the angle between the specimen axis and the (0001) plane of the crystal. From the family of expansion vs temperature curves, the instantaneous expansion coefficients were calculated at 100°C intervals, and Fig. 1 shows a plot of these derived coefficients for each orientation. Coefficients were obtained from the raw data using the chord-slope method. The curves in Fig. 1 indicate a positive temperature coefficient for low x, a negative temperature coefficient for high X, and an expansion coefficient invariant with temperature, amounting to about 43 x 106, for a crystal orientation of approximately 52 deg. From the available data, it is possible to arrive at the values for the axial thermal-expansion coef- ficients a, and a, by extrapolating the curve of the expansion coefficient vs orientation to zero and 90 deg orientations. More precise values for a, and a, are obtained if the instantaneous coefficients are plotted against the square of the cosine of the orientation angle, as shown in Fig. 2. The extrapolated terminal points of the best straight line through the data points of such a plot define the axial coefficients at that temperature. The theoretical justification for the linearity of this curve comes from inspection of a rearranged form of the general anisotropic equation for expansion: ax=a, + (a, -a,)cosZx [1I in which a, is the expansion coefficient in the "c" or [0001] direction and a, the expansion coefficient in the "a" or [ll50] direction. From Fig. 2, the axial coefficients at 100°C, for example, were found to be a, = 13.9 x 10'6 and a, = 62.4 x 10"6 per "C. In the same way, the coefficients were determined for 20", 200°, 300°, and 400°C. This data is summarized in Fig. 3. It is of interest to observe the common intersection of the curves in Fig. 3, which seems to bear out the earlier observation, from Fig. 1, of a

Jan 1, 1964

By A. B. Kinzel

SILICON, unfortunately, is not in the same category as some other metals with respect to the absolute value of the highest purity material prepared. Tucker, in England, and Becket, in this country have devel-oped procedures for the purification of silicon, and Becket's process has made industrially available silicon of a purity of 99.8 per cent or better. However, even the term "silicon metal" may be a paradox to some people, although not more so than the term "amorphous" as applied to silicon. These terms come from commerce, and "metal" is used to distinguish industrially high-purity silicon from ferrosilicon, silicon-aluminum, and other high-silicon alloys. Alloys of iron and silicon with 95 per cent Si or less are known as ferrosilicon, with a designation of silicon content where necessary, and the terns "silicon metal" is invariably applied to the grades containing 96 per cent Si and higher. The Becket method of purification was referred to in a monograph on the Alloys of Iron and Silicon, and the Tucker procedure was described in some detail in that publication.1 In order to get silicon of still higher purity, a sample, prepared by the Becket method, analyzing 99.84 per cent Si, 0.020 Fe, 0.016 Al, 0.005 Ca, 0.025 C, 0.033 per cent 02, 0.006 H2, 0.006 N2, and 0.001 per cent Mn, was crushed to pass a 100-mesh sieve said further purified by a modification of Tucker's procedure. This modification differs from Tucker's method in that after treatment the 100-mesh alloy was centrifuged to remove the last traces of acid. Our directly determined analysis of this purified sample shows: Si, 99.952 per cent; Fe, 0.009; Al, 0.012; Ca, 0.003; C, 0.006; or 99.982 per cent. Silicon by difference equals 99.97 per cent. The variation between silicon by difference and silicon as directly determined is important. Although Tucker mentions the amount of insoluble matter in commercial silicon both before and after his treat-ment, he does not mention the existence of carbon in the alloy; neverthe-less, lie reports silicon by difference. Any final figure as to purity obviously results from agreement between direct determination and

Jan 1, 1939

By J. A. Solomon, H. E. Shafer

As a result of the great demand for metallic aluminum and the scarcity of domestic bauxite, numerous processes have been proposed and developed to extract the oxide of this metal (Al2o3) from clays, shales, and other low-grade raw materials.1-4 An exploratory study was made to extract alumina with sulfuric acid from relatively impure coal-associated draw-slate and to establish an optimum calcination temperature for maximum recovery. Draw-slate is the term used to describe the hard, slaty material directly overlying many coal seams and may range from a few inches to a few feet in thickness. Alumina content of these draw-slates may range from as low as 15% to over 35%. Although the following results are limited in scope, some of the data should prove beneficial for further work. Coal-associated clay, shale, and gob heretofore have been considered nuisance products; any successful alumina extraction process would greatly enhance the worth of these almost limitless resources. This research demonstrated that an optimum calcination temperature for alumina extraction could be established for a coal-associated draw-slate and that this temperature was lower than those suggested in previous work with unspecified clay types. It should be noted that slightly higher acid concentrations and somewhat finer particle sizes were used in this series of tests as compared to previous work. EXPERIMENTAL WORK Essentially, alumina extraction proceeded in four steps: 1) grinding, 2) calcining, 3) acid leaching, and 4) precipitating of aluminum sulfate. An initial series of reference tests to determine the optimum temperature of solubility for alumina was run using relatively pure kaolin. Kaolin forming clays are hydrous aluminum silicates of the approximate composition 2H2O A12O3 2SiO and kaolinite is the mineral that predominates in most kaolins. 5 The kaolin was ground to -30 mesh and subsequently calcined at temperatures ranging from 100°C to 700°C. Each sample was calcined for 1 hr in a muffle furnace and in turn subjected to sulfuric-acid leaching (H2SO4, 9 molar) for a 2-hr period. The filtrates were analyzed for the amount of dissolved alumina recovered.637 Also, each sample of leached or barren residue was

Jan 1, 1968

By T. C. Campbell

Copper is extracted from the Chuquicamata oxide ore by a hydro-metallurgical process. The ore is crushed to 3/8-in. sizing, and leached with a sulfuric acid electrolyte. Chlorine is precipitated. and the ferric iron reduced in the enriched electrolyte, after which the copper is recovered by electrolysis with insoluble Chilex and lead-antimony anodes, the spent electrolyte being returned to leaching. Cathodes are melted and refined in market furnaces and cast into commercial wire-bar and cake shapes. Sulfuric acid for the process is supplied by the brochantite in the ore. Water for washing the ore is advanced through the solution system and finally after cutting by electrolysis to from 6 to 16.0 grams per liter is completely stripped by the cuprous chloride method and run to waste. The cuprous chloride so obtained plus that resulting from precipitation of chlorine from strong solution is dissolved in ferrous chloride brine and the copper cemented on scrap iron. The cement so obtained is in part used to reduce ferric iron in electrolyte, and as it is of exceptional purity is also furnace refined and cast into an exceptional quality of fire-refined copper. Fig. 1 is flow sheet of the operation. The reduction plant is composed of seven divisions as follows: (1) crushing, (2) leaching, (3) tailings disposal, (4) dechloridizing, (5) sulfur dioxide treatment, (6) electrolytic tank house, (7) smelting and melting. A brief description of each of these divisions will be given. Crushing Ore that is frequently blocky and up to 5 ft. in section is crushed in four stages to 3/8-in. mesh. The first stage of crushing is called primary and the next three stages are known as secondary crushing. The first stage is accomplished with two 60-in. Superior McCully gyratory breakers by which the ore is broken to 9 in. The second coarse gyratory breakers crush the ore to 3-in. sizing. There are seven No. 10 McCully gyratories in this service, for one of which recently has been substituted a No. 7 cone crusher. The third and fourth stages of crushing are accomplished by fifty 48-in. disk crushers. In the third stage 14 disk crushers break the ore to 1 in., and in the last stage 36 disk crushers turn out the final product of 0.371-in. sizing. Reference to the flow sheets of Figs. 2 and 3

Jan 1, 1934

By M. A. Jones

Performance of a pilot flood in eastern Kansas indicates improved recovery and accelerated production resulting from mobility ratio control obtained by adding a high moleculur weight polymer to injected water. Ultimate pilot flood recovery of the 75-cp reservoir oil is anticipated to he 350 bbllacre-ft during the six-year flood life. Cumulative recovery to date totals 242 bbl/acre-ft after 30 months of injection. Reservoir parameters and flood performance are disc~rssed in detail to provide a case history relating to the feasibility of polymer flooding. INTRODUCTION The Vernon Polymer flood of the Brazos Oil & Gas Co., located in Sec. 2, 24 S., 16 E., Woodson County, Kans., was begun in late Oct., 1963. Injection of a 500 ppm solution of Pusher* polymer was initiated in a 15-acre pilot to determine feasibility of improving oil recovery by mobility ratio control. Performance to date and the anticipated recovery of 115,000 bbl from the 329 acre-ft pilot indicate efficient displacement of the viscous reservoir oil. Dependence of flood recovery on the displacing fluid-displaced oil mobility ratio has long been recognized. Displacement efficiency of the reservoir oil decreases with increasing mobility ratio. Therefore, economic reduction of this parameter is an attractive approach to improving oil recovery. particularly from viscous oil reservoirs where mobility ratios are often quite high. Waterflood mobility ratios can be controlled by dilute solutions of certain polymers'; which provide a displacing fluid with considerably reduced mobility. Reservoir parameters and flood performance are emphasized in this paper to provide data relating to polymer flood applicability. THEORY The polymer used in the Vernon pilot exhibits the property of unusually high resistance to flow through porous media.'.' A small concentration in injected water effects a mobility reduction considerably beyond that predicted from increases in bulk viscosity as measured in a capillary or rotating cup viscosimeter. Water mobility in the subject flood is reduced by a factor of 7.55 with the addition of 500 ppm Pusher chemical to injected water while the bulk viscosity is increased from 1.03 to only 1.40 cp at 75F. The exact flow mechanism causing resistance to flow has not been established; however, it is observed only in tortuous flow systems and appears to be influenced by brine properties, rock properties and oil saturation. Consequently, mobility reduction is estimated from flow studies for a given sample of reservoir rock and is expressed in terms of a "resistance factor".' This parameter is defined as the ratio of brine mobility to polymer solution mobility measured at residual oil saturation: RF= ?m/?p= (km/µm)/(kp/µp)M.......(1) where RF = resistance factor (dimensionless) k*. and A, = mobility of water and polymer solution, respectively k,,, and k,, = water and polymer permeability, respectively (darcies) µie and µp = viscosity of water and polymer, respectively (poise). Eq. 1 is similar to the expression commonly used to relate the relative mobility of the water bank at residual oil saturation to oil-bank mobility at irreducible water saturation: where M =mobility ratio of water to oil (dimension-less) A, = mobility of oil k, = permeability to oil (darcies) )i* = oil viscosity (poise). From Eqs. 1 and 2 it is apparent that the polymer-solution-to-oil-mobility ratio can be approximated by dividing the brine-to-oil ratio by the resistance factor RF: Mpn=M10/RF...........(3) where M,-, = mobility ratio of polymer solution to oil (dimensionless). This simplified presentation assumes no dilution of the polymer solution by connate water. Also, the presence of a connate water bank separating the polymer solution from the displaced oil is not considered. Displacement of oil by waterflooding is increasingly dominated by viscous forces as the viscosity of the oil displaced increases. When the oil mobility is less than that of the injected water, the oil-water interface is not a piston-like front.3,' Instead, instabilities occur at the flood front in the form of viscous fingers which tend to penetrate the less mobile oil bank. The magnitude of these protrusions, or the degree of channeling, increases

Jan 1, 1967

By R. F. Nielsen, R. G. Bentsen

Jan 1, 1966

By Robert T. Ault

The nature of fracture in unnotched tensile and notched tensile sheet and round specimens and V -notched and precracked Charpy-type sheet specimens of both wrought stress -relieved and re-crystallized molybdenum was investigated over the temperature range of -78° to 300°C. The sharp rise in fracture stress, for unnotched tensile samples, as the temperature is increased above the brittleness transition temperature (Tb), is found to be a result of an increase in flow stress due to plastic constraint and an increased strain rate which results from the onset of necking at Tb . Quasi-brittle fracture in notched tensile samples is found to occur when the tensile stress at the elastic-plastic interface, beneath the root of the notch, reaches a maximum critical value, which is independent of test temperature over the range from —78° to 25°C, but dependent on microstructure. In the temperature range between 150° and 300°C, unnotched tensile and notched tensile samples alike are found to fracture by a ductile fibrous tearing process which is discontinuous in nature, as a result of the competition between the processes of tearing, through continued plastic flow, and local work hardening. Results from the V-notched and fatigue-cracked Charpy-type impact tests demonstrate that crack initiation is the governing factor in the low-temperature (-78° to 25°C) fracture process for molybdenum. In recent years, there has been considerable interest, and a commensurate number of investigations, concerning the duc tile-to-brittle transition in refractory metals. There has not developed, however, a satisfactory explanation for the uniaxial tensile fracture-stress transition which accompanies the well-known ductility transition. This matter therefore warranted investigation. In a similar manner, the increased importance of notched tensile strength values in evaluating materials and in design criteria requires a better understanding of the factors which control the fracture behavior of notched tensile samples. In a previous investigation1 of the nature of initial yielding and fracture in notched sheet molybdenum at room temperature, it was suggested that plastic constraint was the controlling factor in governing the fracture behavior of notched samples. The final portion of this investigation was concerned with the fracture toughness of molybdenum. Of particular interest was the comparison of effective surface energies for fracture in V-notched samples with fatigue-cracked, Charpy-type samples in order to ascertain the relative importance of the initiation and propagation phases of the fracture process. MATERIALS AND TEST PROCEDURE Sheet tensile, notched tensile, and V-notched Charpy specimens were prepared from 50-mil, stress-relieved molybdenum sheet.* Half of these specimens were vacuum-annealed for 1 hr at 1200°C and 7.5 x l0-5 Torr and furnace-cooled to produce a relatively uniform grain size of 0.30 mm diameter. Round notched and unnotched tensile specimens were prepared from warm-rolled and swaged 5/8-in.-diam bar.* These specimens were vacuum-annealed for 1 hr at 1350°C and 5 x lom5 Torr and furnace-cooled to produce a grain diameter of 0.11 mm. Material analyses in ppm were: C N O H Sheet 290 10 4 1 Bar 50 30 30 2 The unnotched tensile sheet specimens had a 1/4-in. gage width and a 1-in. gage length. The unnotched tensile round specimens had a 9/32-in. gage diameter and a 1-1/4-in. gage length. The sheet and round notch tensile specimens are shown in Fig. 1. All of the notched tensile specimens tested were tandem-notched in order to study the location and mode of fracture initiation. The tandem-notched specimens were carefully machined so that the largest variation in notch depth on a single specimen was 0.001 in. Thus, when fracture occurs, the extent of plastic deformation that exists in the un-fractured notch section is that which exists just prior to fracture. All of the tandem-notched sheet specimens were electrolytic ally polished and chemically etched prior to testing. The V-notched Charpy-type specimens machined from the 50-mil sheet material had the standard dimensions of 2.125 in. long, 0.394 in. deep, and a 0.010-in. root radius. The unnotched tensile and notched tensile tests were conducted over the temperature range of -78° to 300°C, in a 10,000-lb Instron Universal Testing Machine, at a constant crosshead speed of 0.020 in. per min. The V-notched Charpy-type

Jan 1, 1964

By C. L. Sollenbeger, R. B. Greenwalt

PERHAPS the most widely used dispersants or gangue depressants in nonmetallic flotation are sodium silicates, which vary in silica-to-soda ratio from 1 to 3.75. Typical manufactured silicates in order of decreasing solubility and increasing amounts of silica are Metso, silica-to-soda ratio of 1.00; D, 2.00; RU, 2.40; K, 2.90; N, 3.22; and S-35, 3.75.* References in flotation literature1,2 to the use of sodium silicates are often weak because they fail to mention the type of silicate used. Metso and silicate N have occasionally been mentioned, but when the type of silicate is not mentioned, it is usually assumed to be N, the cheapest of the soluble silicates and the one recommended by sodium silicate manufacturers as a flotation agent. In the All is-Chalmers Research Laboratories a systematic study was made of the effect of different alkali-silica ratios on the concentration by flotation of two scheelite ores. One of these was a high grade ore from the Sang Dong mine in Korea. The effect of such factors as pH; addition agents; and conditioning time, temperature, and pulp density on the flotation efficiency of this ore have been described previously. The other ore was a low grade ore from Getchell Mines Inc., Nevada. The mineralogy and techniques of concentrating this ore have been described by Kunze. Hereafter these ores will be referred to as the Korean and Nevada ores. Experiments were made with both to determine the effect of three factors—-type of silicate, concentration of silicate, and pH of the pulp—on recovery and grade of tungsten in a rougher concentrate. Average WO, content of the Korean ore was 1.50 pct and of the Nevada ore 0.27 pct. The predominant tungsten mineral in both ores was scheelite, which was accompanied by a small amount of powellite. The powellite and scheelite were finely disseminated through both ores and required a —200 mesh grind for liberation. Major gangue minerals in the Korean ore, in decreasing order of abundance, were amphi-boles, quartz, biotite, garnet, fluorite, and calcite. Bulk sulfides composed about 3 pct of the total weight. Gangue in the Nevada ore, in descending order of abundance, was garnet, alpha quartz, calcite, phlogopite, wollastonite, and amphiboles. Sulfide minerals were 3 to 4 pct of total weight. Batch flotation experiments were made with 500-g samples of ore, each sample wet-ground to 90 pct passing 200 mesh. The finely ground ore was floated in a Fagergren batch cell at 25 pct solids. The natural pH of the Nevada ore was 8.9 and of the Korean ore, 8.5. The D, RU, K, N, and S-35 sodium silicates were obtained in colloidal dispersions with varying amounts of water. The most alkaline, Metso, was in dry powdered form. For convenience in addition, 5 pct solutions by weight were prepared from each of the silicates, on the basis of dry sodium silicate dissolved in the correct amount of distilled water. Chemical analyses of the various silicates are given in Table I, together with the pH of the 5 pct solutions. A preliminary bulk sulfide float was made with secondary butyl xanthate as the collector and pine oil as the frother. The WO] analysis of the sulfide concentrate was nearly 1 pct for the Korean ore and about 0.1 pct for the Nevada ore. The tungsten contained in the sulfide concentrate constituted about 3 pct of the total tungsten in each ore. No effort was made to recover these tungsten values. The scheelite was floated with oleic acid. Adjustments in pH were made with sulfuric acid or sodium carbonate. A 1 pct solution of 85 pct Aerosol OT was sprayed on the froth and sides of the cell during the scheelite float to aid in dispersing the minerals and to decrease the entrapment of gangue particles. Six tests were planned for each of the six types of silicate in which concentrations of 1, 2, and 4 1b of silicate per ton of dry ore were investigated at both 6.5 and 10 pH. All tests were made at room temperature. The performance of each silicate was judged from the grade and recovery of WO, in the scheelite rougher concentrate. Tungsten recovery was calculated on the basis of the scheelite remaining in the ore after the preliminary sulfide float. Testing of each silicate at three levels of concentration and two levels of pH required 36 tests with each scheelite ore. Variance analyses were performed on the concentrate grades and recoveries to determine whether or not the type of sodium silicate, the concentration of sodium silicate, or the pH significantly affected recovery or grade. Results Concentrate Grade: A variance analysis of the concentrate grades for the Korean ore showed that concentration of the silicate and pH of the ore pulp were major factors in producing a high grade concentrate. Also, the silica- to-so da ratio was important as an interaction with pH. The concentrate grade vs silica-to-soda ratio is plotted in Fig. 1. The curves show that the concentrate grade improved with an increase in concentration of sodium silicate and also

Jan 1, 1959

By B. A. Eaton

The subject of many discussions and technical papers in the last 20 years has been the prediction of the well-bore pressure gradients that are required to induce or extend fractures in subsurface formations. The subject merits this attention because of the frequently recurring problems that arise from an inability to predict fracture pressure gradients. Encountered in several common types of operations in the oil industry are problems associated with the prediction of formation fracture pressure gradients. When wells are being drilled in both new and old fields, lost circulation is often a very troublesome and expensive problem. Complete loss of circulation has been disastrous in some cases. Many times, such disasters could have been avoided if techniques for calculating fracture pressure gradient had been employed in the well plans, and if casing strings had been set, and mud weight plans had been followed accordingly. In areas of abnormally pressured formations, the prediction of fracture gradients during the well-planning stage is extremely important. In fact, it is as important as the prediction of formation pressure gradients, which has received a great deal of attention in recent years. There are several published methods used to determine fracture pressure gradients. However, none of these methods appears to be general enough to be used with much reliability in all areas. In 1957, Hub-bert and Willis published a classical paper that included the development of an equation used to predict the fracture extension pressure gradient in areas of incipient normal fau1ting.l Overburden stress gradient, formation pore pressure gradient and Poisson's ratio of rocks were the independent variables that were shown to control fracture pressure gradient, the dependent variable. In 1967, Matthews and Kelly published another fracture pressure gradient equation that is different from that of Hubbert and Willis in that a variable "matrix stress coefficient" concept was utilized.3 Later the same year, Costley wrote about a similar idea.5 Goldsmith and Wilson used a least-squares curve-fitting technique and field data from the Gulf Coast area to correlate fracture pressure gradient with formation pore pressure gradient and formation depth.' They noted that the fracture pressure gradient increased with increasing depth while the pore pressure gradient remained constant. In each of these cases, the problem for which a solution was sought was to determine the bottom-hole pressure gradient required to initiate or extend a fracture. Results of the previous work show that fracture pressure gradient is a function primarily of overburden stress gradient, pore pressure gradient, and the ratio of horizontal to vertical stress. There is argument for a fourth variable in that in many cases breakdown fracture pressure gradient is greater than the fracture extension pressure gradient. However, if the fracturing fluid is able to penetrate the formation through the pores or existing cracks, there is very little

Jan 1, 1970

By K. C. Ko, D. A. Pifer

The theory of probability with respect to the failure of structures is discussed. It is shown that the probability of safety, probability of failure, and factor of safety are directly related to each other and that as standard deviation approaches zero, the probability curves approach a unit step function which can be interpreted as the conventional factor of safety. The factor of safety is numerically the ratio of the load to the strength of the material at a particular point or plane in an engineering structure. Mathematically, it can be illustrated as a unit step function undefined at the point where the factor of safety is one (Fig. 1). The uncertainty of variation of soil and rock strengths require factors of safety much higher than those used in other engineering structures. The factors of safety used are generally determined from past experiences with similar structures. However, the factors thus determined do not have any theoretical meaning or do not indicate any relationship to actual safety.' It seems, therefore, logical to develop a system with a sound theoretical basis to obtain a more suitable factor of safety for structures. An application of elementary theory of probability indicates that it is possible to relate this number to theoretical and practical engineering facts.

Jan 1, 1971

By E. L. Tucker

PREVIOUS investigations of E. L. Tucker and R. E. Head' related in particular to the effect of cyanogen compounds on galena, sphalerite, and pyrite, and their behavior in the presence of such com- pounds. The present contribution deals with the float- ability of the copper minerals and of pyrite. All experiments were performed on minus 200-mesh material, the method of preparation being the same as that described in the previous article. Flotation tests were made in a small mechanically agitated machine of 50-g. capacity designed in this laboratory by Gates and Jacobsen." General Naval Stores pine oil No. 5 and potassium xanthate were used in constant amount in all of the experiments; these agents were selected because of their present wide use in the flotation of copper minerals. Lime additions were varied to secure four 'final hydrogen ion concentrations, namely, neutral, 0.0005 per cent CaO, 0.003 per cent CaO, and 0.02 per cent CaO. The variation in lime addition was introduced to include the range of alkalinities employed in actual flotation practice.

Jan 1, 1926

?AARONSON, ALFRED E , Vice-Pres , Mid-Co Petroleum Co, Mid-Co Bldg, Tulsa, Okla '18 ABADIE, EMILE R, Min Engr Address wanted '76 || ABADILLA, QUIRICO A, Geol Dept, Cia Mexicana de Petroleo "El Aguila," S A, Edifico "El Aguila," Esq, Calles Aurora y Comercio, Tampico, Tamaulipas, Mexico '20 ABBOTT, CLARENCE E, Gen'l Supt, Iron Mines Div, Tenn Coal, Iron & R R Co, 1405 Minnesota Ave, Bessemer, Ala '04 ABBOTT, LOUIS BENJAMIN, Chief Engr, Consolidation Coal Co, Elkhorn Div, Jenkins, Ky '15 ABBOTT, ROBERT R Met Engr, Peerless Motor Car Co, Cleveland, Ohio '12 ABE, ASAKA Fujita Min Co, Dojima, Osaka, Japan '14 ABEEL, GEORGE H, JR, Cons Min Engr, 635 St Paul Ave, Los Angeles, Cal '18 ABEL, ALBERT E, Asst to Pres, Valley Mould & Iron Corpn, Sharpsville, Pea '19 ABEL, W D, Min Engr Forum Bldg, Sacramento, Cal '19 ABELL, O J, Pres & Treas, Abell-Howe Co, 332 South Michigan Ave, Chicago, Ill '19 ABERNATHY G E Willow Springs, Mo '20

Jan 1, 1923

Following is the personnel of the various society and national committees on which the A I M E is officially represented United Engineering Society Officers of the Board, 1921 JOHN VIPOND DAVIES, President CALVERT TOWNLEY, 1st Vice-president ALFRED D FLINN, Secretary WILLIAM L SAUNDERS, 2d Vice-president JOSEPH STRUTHERS, Treasurer Finance Committee House Committee CALVERT TOWNLEY, Chairman ALFRED D FLINN, Chairman GEORGE H PEGRAM . H S CROCKER WILLIAM L SAUNDERS BRADLEY STOUGHTON GEORGE M BASFORD CALVIN W RICE F L HUTCHINSON

Jan 1, 1923

By David A. Thomas, David N. French

The lrnrdness of WC crystals has been measured with the Knoop indenter at loads of 100 and 500 g on the (0001) and (1070) planes. The hardness as tneasitred on the basal plane is 2400 kg per sq mm and shows little anisotropy. The hardness on the prism plane, however, shows a marked orientation dependence, with a low value of 1000 kg -per sq mm when the long axis of the Knoop indenter is oriented parallel to the c axis and a high value of 2400 kg per sq mm when the indenter is perpendicular to the c axis. Slip lines (Ire observed surrounding the microhardness indentations and they show slip on (1010) planes, probably in [0001] and (1120) directions. This slip behavior can be explained by the crystal structure of TVC, which is simple hexagonal with a c/a ralio of 0.976. The hardness anisotropy call be explained by [0001]{1010} and (1130) {10l0] slii) and the resolved shear-stress analysis of Daniels and Dunn. HARDNESS anisotropy is a well-known phenomenon and has been reported for many metals, with both cubic and hexagonal structure.1-6 For hexagonal tungsten carbide, WC, a wide range of hardness values is reported in the literature. For example, Schwarzkopf and Kieffer7 give a hardness of 2400 kg per sq mm and report a value of 2500 kg per sq mm by Hinnüber. Foster and coworkerss give the average Knoop microhardness as 1307 kg per sq mm with a maximum value of 2105 kg per sq mm. Although these values and the structure of WC suggest the likelihood of hardness anisotropy, no such measurements have been made. We first detected a large apparent hardness anisotropy in WC crystals about 75 p large, in over-sintered cemented tungsten carbide. Prominent slip lines also occurred around many indentations. This report presents further observations and interpretations of hardness anisotropy and slip in WC crystals obtained from Kennametal, Inc. Both Knoop and diamond pyramid indenters were used on a Wilson microhardness tester with loads of 100 and 500 g. EXPERIMENTAL RESULTS The carbide crystals tended to be triangular plates parallel to the (0001) basal plane of the hexagonal structure. The side faces were parallel to the ( 1010) prism planes. Specimens were mounted approximately parallel to these two types of faces and metallographically polished. Laue back-reflection X-ray patterns were used to orient the specimens, which werethen ground to within ±1 deg of the (0001) and (1010) planes. The Knoop hardness values measured on the basal plane are plotted in Fig. 1. There is only a small anisotropy, with a hardness range of 2240 to 2510 kg per sq mm. The additional points at angles from 52.5 to 67.5 deg confirm the sharp minimum hardness at 60-deg intervals, consistent with the sixfold hexagonal symmetry. The average hardness of all values obtained on the basal plane is 2400 kg per sq mm. While the basal plane shows only slight anisotropy, the (1010) plane exhibits marked hardness anisotropy, from 1000 to 2400 kg per sq mm. Fig. 2 shows the hardness as a function of the angle between the long axis of the indenter and the hexagonal c axis, the [0001] direction. The minimum and maximum occur when the indenter is oriented parallel and perpendicular to the [0001] direction, respectively. The anisotropy of the prism plane is contrary to that reported for hexagonal zinc and hard- However, the basal-plane anisotropy is similar to these two metals.1'2 To check the accuracy and reproducibility of the measurements, a series of fifteen impressions was made at 100-g load in the same orientation in the same area of the specimen surface. The average for all was 2040 kg per sq mm, with a range of 1950 to 2130 kg per sq mm, giving an accuracy of about ± 5 pct. Thus the slight anisotropy on the basal plane is almost within experimental error. Fig. 3 shows slip lines around the Knoop indentations on the basal plane. The slip traces are in directions of the type (1130). The presence of slip steps on the basal plane indicates that the slip direction lies out of the (0001) plane. Because WC has a c/a ratio of 0.976,' the shortest slip vector is [0001], which suggests slip systems of the type [0001] (1010). Fig. 4 shows slip lines around the Knoop intentations on the (1010) plane. These slip lines are inconsistent with [0001] slip but can be

Jan 1, 1965

By R. I. Jaffee, H. R. Ogden, D. J. Maykuth, W. L. Finlay

Titanium with up to 7.5 pct Al forms single-phase a alloys that are work hardening, not heat treatable, and ductile as welded. The high aluminum Y phase alloys are not usefully ductile, despite their low hardness, but they have excellent oxidation resistance and hot hardness. LITTLE has been published on the effect of aluminum on the mechanical properties of titanium, despite the fact that two of the first four commercial alloys released by titanium alloy producers contained aluminum as an important alloying ingredient. These were both a-ß type alloys, one containing 4 pct A1 and 4 pct Mn and the other containing 3 pct A1 and 5 pct Cr. However, neither of these alloys reflects the true alloying nature of aluminum, since each contains important amounts of the ß-stabilizing addition, manganese or chromium. In a previous publication on the constitution of Ti-A1 alloys,' the authors showed that aluminum had a high solubility in a titanium and was strongly a stabilizing in nature. About 25 wt pct A1 is soluble in a titanium at room temperature. With as little as 5 pct added aluminum, the a transus temperature is increased from 882" to 970°C, and the correspond- ing ß transus to 995°C. The first intermediate phase in the Ti-A1 system is the ordered face-centered tetragonal TiAl phase, which extends from 34 to 46 wt pct Al. The present paper presents hardness data over the entire titanium-rich region of the system, through the TiAl phase, and mechanical property data on the alloy range capable of being fabricated. Methods of Investigation The materials used and general preparation of alloys, as well as experimental techniques used in the investigation of Ti-A1 alloys, have been adequately described.' Briefly, the preparation consisted of arc-melting the desired amount of high purity aluminum (99.99 pct purity) and iodide titanium (maximum hardness 100 Vickers, as deposited) together on a water-cooled copper hearth under a partial pressure of argon. The resulting button ingot was of such form that it could be fabricated without further treatment. All work was done using ingots weighing 10 to 15 g, which were hot-rolled to strip form whenever possible. Tensile samples were 3.25 in. long, 0.5 in. wide, and 0.040 in. thick, with a reduced section 1.25 in. long by 0.250 in. wide. The gage length used was 1 in. Electrical strain gages mounted on the

Jan 1, 1954

By Martin Rubenstein

CdS crystals have been grown from a number of metallic solvents such as bismuth, tin, lead, and cadmium. Etching studies have shown that plastic deformation occurs if the crystals are not removed from the solvent prior to the solidification of the solvent, on cooling. The deformed crystals show a umique exciton fluorescence as a function of edge dislocation density. If one grows the CdS in the eutectic alloy of the above four metals (commonly called Wood's metal) the crystals can be removed from the solvent with hot water and no plastic deformation occurs. In this paper, the liquidus solubility measurements of CdS, as a function of temperature, are presented. The data were obtained using a high -temperature filtration technique. CADMIUM-SULFIDE crystals have been grown from a number of metallic solvents1 such as cadmium, bismuth, tin, and lead. Liquidus solubilities of CdS in cadmium,2 bismuth,3 and tin4 have already been measured. Crystals of CdS, in all four metals, have been grown by solution growth: 1) by cooling a saturated solution and 2) by a solution transport method.1'"1 CdS crystals grown in these four solvents have a few characteristics in common: 1) 1.8°K photolumines-cent emission consisted mainly of the radiative recombination of the bound exciton commonly known as I,, 2) slip lines which could easily be seen by the naked eye, and 3) edge dislocation densities in the order of l05 per cu cm.1 It was decided that these slip lines and the high edge dislocation densities were caused by a plastic deformation of the CdS crystals. It was felt that this plastic deformation did not occur during the growth of the crystals nor during the cooling of the solution, but did occur when the solvent which was in contact with the crystals froze. If these assumptions were valid, the slip lines and the high number of dislocations could be reduced or eliminated by removing the crystals from the solvent before the solvent froze. Since crystals of CdS had already been grown separately in such solvents as bismuth, lead, tin, and cadmium, it was felt that crystals could be grown in a eutectic mixture of these four metals. In this work a eutectic (or near eutectic) mixture of bismuth, lead, tin and cadmium in the proportion 50, 26.5, 13.5, and 10 wt pct, respectively, was used to grow CdS crystals. Such a mixture has a melting point of about 70°C and is close in composition to the alloy commonly known as Wood's metals. If the crystals could be grown from this mixture of solvents, and if hot water (>75°C) could be used to separate the crystals of CdS from the metallic solvent, it was hoped that CdS crystals could be grown with little or no plastic deformation which had been ob- served when crystals were grown from these solvents uncombined. CdS crystals were grown from this low melting eutectic mixture of bismuth, lead, tin, and cadmium using the solvent transport method. CdS powder and the appropriate amount of metals were sealed in a quartz tube under a pressure of about 5 X 10-6 torr. This ampule was then placed in a vertical position in a furnace. The temperature was raised to about 900°C. The furnace was designed so that the top of the liquid column within the ampule was between 10° to 40°C higher than the bottom of the liquid column. These temperatures were measured on the outside of the quartz ampule. The ampule was maintained at temperature for 7 to 14 days (depending on the temperature at which transport was taking place) and then the furnace temperature was lowered until the temperature was about 125°C. The ampule was then removed from the furnace, placed in water maintained at about 90°C, and opened in this 90°C environment. The crystals could then be removed from this two-phase liquid (Wood's metal and water) by mechanically picking them out. Alternatively, the crystals could be quantitatively removed by adding an excess of mercury to the mixture of metals, crystals, and hot water. The hot solution of metals and the hot water could be evacuated using a small diameter tube connected to a vacuum. Small amounts of mercury and water could be removed by heating the crystals in vacuum. Crystals prepared using this technique showed no evidence of slip. However, some of these crystals did show edge dislocation densities as high as l04 per cu cm. Some few selected crystals showed no dislocations. Single crystals of CdS were grown as large as 5 by 5 by 0.5 mm. The ampules for the growth of these crystals were 13 mm O.D., 11 mm I.D., 150 mm! LIQUIDUS SOLUBILITY MEASUREMENTS The CdS starting materials was G.E. 118-8-2 powder which was fired in H2S at 1000°C, and then a vapor transport technique5 was applied to produce a "sound" mass of CdS. The Wood's metal was prepared by weighing out bismuth, lead, tin, and cadmium in the proportions of 50, 26.5, 13.5, and 10 wt pct, respectively. The bismuth, cadmium, and lead were from the American Smelting and Refining Co. (ASARCO) and all had purities of 99.999+ pct. The tin was 99.9999 pct spectroscopic grade from the Vulcan Materials Co. The appropriate mixture was placed in a quartz tube, evacuated to a pressure of 5 X 10-6 torr, melted to a liquid, cooled to room temperature under this same vacuum. This ingot was then placed in another quartz tube, evacuated to 5 x l0-6 torr, and sealed off under vacuum. The ampule was then horizontally placed in a furnace. The temperature was raised to 600°C, and over a period of several hours the ampule was vigorously shaken several times. The ampule was then removed from the furnace, and the metallic liquid was

Jan 1, 1970

IRON AND STEEL COMMITTEE ALBERT SAUVEUR, Chairman .A. A. STEVENSON, Vice-Chairman HERBERT M. BOYLSTON, Secretary, Abbot Bldg., Cambridge, Mass. Guilliaem Aertsen, Robert W. Hunt, .-E: Gybbon Spilsbury, John Birkinbine, J. E. Johnson, Jr., J. S. Unger, William H. Blauvelt, William Kelly, Felix A. Vogel, Charles A. Buck, Charles Kirchhoff, Leonard Waldo, James Gayley, Richard Moldenke, William R. Walker, Henry D. Hibbard, Joseph W. Richards, William R. Webster, Henry M. Howe, C. F. W. Rys, Frederick W. Wood. Manuscripts of the following papers on iron and steel for the February -meeting have been received and accepted: Albert Sauveur, Notes on Heating and Cooling Curves of Professor Carpenter's Electrolytic Iron. H. M. Howe and A. G. Levy, Notes; on the Plastic Deformation of Steel during Overstrain.

Jan 2, 1914

By Charles Legrand

The power plant of the Burro Mountain Copper Co. is located near Tyrone, N. M., at 5,950 ft. elevation. It is interesting because it uses the largest stationary Diesel engines in the United States. The general layout of the plant is shown in Figs. 1 and 2. The building is of steel-frame construction with hollow-tile walls and corrugated-iron roof; the floor is concrete and all windows have steel frames, making a fireproof building. The building is made to accommodate one more electric generator unit which is being installed; also two 4,000-cu. ft. air compressors direct-connected to three-cylinder engines with the same size cylinders as the electric units. The power plant consists of two 815-kva., 60-cycle, three-phase, 6,600-volt, 180-r.p.m. generators, direct-connected to five-cylinder, two-cycle Diesel engines; the necessary circulating pumps for cooling water, cooling towers and oil storage tanks. A portion of the power is used in the same building as the generating units to run one of two 200-kw. rotary converters delivering 260-volt direct current, and a 2,500-cu. ft. air compressor, delivering air at 95 lb. pressure, direct-connected to a 400-hp., 6,600-volt, three-phase synchronous motor. The remainder of the power is used to run a 1,500-ton concentrating mill 3 ½ miles from the power house, various hoists, pumps and motors about the mines and the lighting and water-works of the town of Tyrone. The direct current is used for electric traction in the mine, also to run hoists and motors that were installed before the new power plant was built. Due to the large hoists that take starting peaks of 250 kw. and the 10-ton electric locomotive that takes at times 150 kw., the load is extremely variable, as shown in. Fig. 3.

Jan 9, 1916

The Inspiration ore body was discovered in 1904, but a 50 -ton mill which was installed to treat the disseminated copper ore promptly failed. The original Inspiration Company then gave way to another, backed by W. B. Thompson, which found 40 million tons of ore. Control then passed to Anaconda and Dr. Louis Ricketts took charge as consulting engineer. He abandoned the mill which had been partially built (to the horror of the stockholders, according to Ira Joralemon) and spent a year and a million dollars experimenting with flotation. The first large copper flotation plant was then built and recovered 50% more copper than had previous gravity mills. It, of course, was highly successful and became the forerunner of modern flotation plants. Inspiration is distinctive as the first fully integrated copper operation in the United States. Three open pits supply sulfide ore to the concentrator and mixed oxide and sulfide ore to a vat leaching and electrowinning operation with leached ore also being sent to the concentrator. In addition, dump leaching supplies pregnant solution for scrap iron precipitation. Inspiration has recently constructed an electric furnace smelter and also operates a continuous casting and rod plant. The new primary 54 in. by 74 in. crusher was completed in May of 1972 and replaced an older unit which was in the way of the open pit development. It is of an unusual circular silo design, which permits dumping 0 from two positions at 90 from each other. One dumping position is in line with the spider, while the other one is at right angles to it. A box girder splitter bar hung from the dump pocket slopes downward from the second dump point, across the spider cap, and rests on the pocket bottom. This splitter was not installed to protect the spider but was to aid in distributing wet ore being dumped from that position to both sides of the spider. Any large amount of wet ore delivered to one side of the spider would pack in the crusher cavity. The round front wall of the 18 ft (5.5 m) deep pocket is unlined except for light steel protection at the top, while the straight back wall is lined with vertically positioned railroad rails. Water for dust suppression is sprayed into the pocket through a cluster of 3/8 in. (9.5 mm) nozzles at the back of the pocket. About 5 gpm (19 lpm) of water are mixed with 90 psi (620 kPa) air upstream of the nozzles to produce a fog which is effective in dust control. The 54 in. Allis -Chalmers gyratory is placed mostly within one quadrant of the circular enclosure to permit placement of the apron feeder on the lowest level in the proper position to transfer crushed ore to the

Jan 1, 1978

By E. A. Gulbransen, A. Pebler

Pressure-composition isotherms have been determinedfor the systems ZrMo2 -H2 between 0" and 900°C at hydrogen pressures between 10-4 and 760 Torr. Tkese studies plus X-ray diffraction analyses of selected compositions show that these intermetallic compounds form a complete series of hydrogen solid solutions within this temperature and pressure range. The partial molar enthalpies and entropies of solution of hydrogen in the three intermetallic compounds were calculated. The solubility of hydrogen in these Laves phases can be related to their free electron concentration. ThE reaction of hydrogen with inter metallic compounds of zirconium with Laves phase-type structure has both theoretical and practical interest. The in-termetallic compounds have structures different from any of the pure metals. In addition the electron to atom ratio of the intermetallics varies appreciably. In the nuclear materials field, these materials are of interest since intermetallic compounds may exist as precipitates in commercial zirconium alloys with transition metals. From studies on the hydrogen reaction, one can establish the stability of the intermetallic compounds relative to the hydrides of the component metals. The intermetallic compounds may also be used as getters for hydrogen and as materials for hydrogen storage. This paper will present the results of equilibrium and structural investigations of the ZrCr2-H2, the ZrV2-H2, and the ZrMo2-H2 systems for the temperature range of 0" to 900°C and the pressure range of Torr to 1 atm. A complete thermochemical analysis will be made of the measurements. Some of the results have been referred to in a survey paper' covering the types of hydrogen behavior observed in zirconium alloys. LITERATURE ZrCr2 is the only intermetallic phase in the Zr-Cr system.' It exists in two allotropic modifications. Between 900" and 1000°C the hexagonal MgZa (C14) structure transforms into the cubic MgCu2 (C15) structure. Beck4 has previously shown that the intermetallic ZrCrz takes up hydrogen into solid solution up to a composition H/ZrCr2 = 1.14 at room temperature and 1 atm hydrogen. A ZrCr2 hydride was stated to have a fcc structure. The binary Zr-V system has been investigated by williams5 although Rostoker and Yamamoto 6 had given earlier a partial phase diagram. ZrV2 was the only intermediate phase found in the system. wallbaum7 reported earlier that the compound ZrV2 hp the MgZm (CJ4) type structure with a = 5.288A and c = 8.664A. This was not comfirmed by Elliot8 nor by Matthias, Compton, and corenzwit9 who reported that ZrV2 has a cubic MgCua LC15) type structure. The latter authors gave a = 7.429A. Also, several authors do not agree on the temperature of the peri-tectic horizontal. It appears that ZrV2 may exist in several modifications. A high capacity for occluding hydrogen was reported by Beck4 who found H/ZrV2 = 1.38 at room temperature and 1 atm of hydrogen. ZrMoz is the only intermetallic phase in the Zr-Mo system.10 ZrMoa is formed by a peritectic reaction of the melt with a molybdenum-rich solid solution. ZrMo2 has the C15-type structure with a = 7.59A. MATERIALS High-purity chromium, vanadium, and molybdenum were used together with grade 1 crystal bar zirconium to prepare the alloys. Table I lists the chemical analysis and the sources for the metals. Ten-gram alloy compacts were levitated and melted in an inert argon

Jan 1, 1968