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By T. S. Lundy, J. I. Federer

Chemically purified Zr95and Cb95 have been used in determining self-diffusion coefficients in the bcc phase of iodide zirconium over the temperature range of 900o to 1750°C. The temperature dependence of the diffusion coefficients could not be described by the usual Arrhenius-type equation over the entire temperature range. Values of the apparent frequency factors (D0)a and activation energies (Q)A for ZR95 varied from 4.8 x 10-6 to 2.5 x 10-2 sq cm per sec and 20,700 to 46,900 cal per mole, respectively, between 900°and 1750°C. Similarly, for Cb95, values of (Do) a and (Q)A varied from about 2.4 x 10'5 to 0.25 sq cm per sec and 26,500 to 56,800 cal per mole, respectively. The data may be described in terms of the a to ß transition temperature of 1136 °K and the absolute temperature of the diffusion anneals by the following eauations: SEVERAL studies on the diffusion of Zr95 in zirconium have been reported.1"5 The values of the constants contained in the usual Arrhenius-type equation for describing the temperature dependence of the diffusion coefficient, D = Do exp [- Q/(RT)], are summarized in Table I. The reported values of Do, the frequency factor, and Q, the activation energy, are seen to vary by factors of about 60 and 1.6, respectively. Examination of the data from any one study indicated reasonable consistency of measurements. The large discrepancies of over-all results, therefore, cannot be explained on the basis of experimental scatter. One possible explanation for the differences in results could be a varying impurity content of the zirconium. Another possibility is that the radioactive daughter, Cb95, influenced the values of some of the determined diffusion coefficients. Therefore, the first of the three purposes of this study was to reconcile the large differences in diffusion results that have been obtained. Several empirical relations have been proposed to relate the diffusion coefficients in solids with physical properties, such as the absolute melting point, the heat of fusion, and the heat of sublimation.

Jan 1, 1963

By Robert Baker

IT has been demonstrated in the laboratory by A. J. Phillips1 that a transformation or conversion from beta directly to alpha may take place in a brass of 61 to 62.5 per cent copper content. The completion of the transformation is dependent on a rapid rate of cooling from the all-beta field through the alpha + beta range to the alpha field. Phillips states that in a slowly cooled alloy containing 62 per cent copper and 38 per cent zinc the copper content of the beta is progressively decreased from 62 to 56 per cent by precipitation of alpha. Rapid cooling tends to prevent the separation of alpha with a copper content of more than 62 per cent by reducing to a minimum the time required for passing through the alpha + beta range. Upon crossing beyond the (alpha + beta.)/(alpha) boundary to the all-alpha region below, the material is transformed from beta to alpha. It is essential that there be no change in composition. Phillips obtained his results by quenching a 62:38 copper-zinc alloy from 850° C. in iced brine at -8° C. No reference to the occurrence of this transformation in a manufacturing process has been found in the literature. An investigation of the microstructures of a number of hot forgings made from leaded brass rods yielded one forging in which this striking conversion from beta to alpha had occurred. The purpose of this paper is to describe the transformation and certain conditions under which it may occur during the hot pressing of brass. The conversion was noted in a shape forged from a leaded brass rod, the composition of which was: Cu, 60.26 per cent; Zn, 37.86; Pb, 1.88. In this study the lead content was disregarded and the alloy was considered to be equivalent to an alloy of 61.40 per cent copper and 38.60 per cent zinc (98.12 parts = 100 parts). The possibility that the lead might obstruct or accelerate the precipitation of alpha was not studied. The lead appears as black globules in the photomicrographs and has a negligible solubility in the alloy.

Jan 1, 1931

By M. R. J. Wyllie, Michael A. Torcaso

With the continued deep drilling of today, increasing numbers of high pressure and high temperature gas-condensate reservoirs are being discovered. Correspondingly, the ranges of properties of gas-condensate reservoir fluids are also being increased. The literature1*'." contains several excellent reviews of the property ranges in presently known condensate reservoirs. When one examines these property ranges, it becomes apparent that the need for reliable data is great if the engineer is to evaluate properly or appraise a gas-condensate reservoir. It is the purpose of this paper to present phase-behavior data that will point out some of the important variables that influence recovery from gas-condensate reservoirs produced by pressure depletion. All these variables may be resolved into two major categories: the hydroc:arbon composition of the reservoir fluid and the physical properties of the reservoir. But the variables which we will consider are those which are usually determined from well tests shortly after discovery and are, consequently, readily available to the engineer. They are as follows: (1) the producing separator gas-stock-tank liquid ratio, (2) stock-tank liquid gravity, and (3) reservoir temperature. Laboratory depletion studies have been performed on several reservoir fluids to illustrate the effect of changes in the cronsidered variables upon ultimate recoveries. The deviation factors, Z, of a series of reservoir fluids were determined experimentally to illustrate their order of magnitude and their effect upon the calculation of in-place reserves existing at high pressurcs and temperatures. The procedures followed in conjunction with the performance of these studies and their associated calculations are illustrated in Fig. 1. For the purpose of this paper, the depletion recoveries are expressed only in terms of pcr cent of initial stock-tank liquid, with the abandonment pressure assumed to be 500 psig. VARIATION OF RECOVERY WITH INITIAL PRODUCING SEPARATOR GAS-STOCK-TANK LIQUID RATIO The producing separator gas-stock-tank liquid ratio has long been one of the principal criteria in assumptions and generalized correlations' used for the estimation of ultimate recovery from gas-condensate reservoirs. Fig. 2 presents a plot of the cumulative recovery of stock-tank liquid at abandonment as a function of the initial producing separator gas-stock-tank liquid ratio. The graphical illustration represents the results of 44 laboratory depletion studies performed on actual gas-condensate fluids. These data are presented not only to illustrate the general trend of ultimate recovery increase with initial producing ratio increase but to emphasize the considerable variation in possible recovery at any given producing ratio. Table 1 presents the properties of two actual reservoir fluids chosen to illustrate the approximate extremities of producing ratios for gas-condensate systems. The fluids designated as A and B exhibited initial producing ratios of 3,400 and 114,000 scf/bbl, respectively. Fig. 3 presents a graphical comparison of the cumulative stock-tank liquid recoveries, calculated from the well-stream compositions and produced volumes for each fluid during depletion. It is to be noted that Fluid A exhibited a recovery of only 16.5 per cent when produced to an assumed abandonment pressure of 500 psig while the corresponding recovery for Fluid B was calculated to be 62.2 per cent. This difference in recovery, which exemplifies the general trend of recovery increase with initial producing ratio increase, is readily explained by consid-

Map

Jan 1, 1939

This is a most interesting paper and some data in it are corroborated by work I did at Watertown Arsenal some time ago with iron powder. I would like to ask Dr. Wulff to say just a little more about the compressive and tensile stresses on the surface of compacts. JOHN WULFF (author's reply).-Our work has not yet been completed; although enough data have been obtained with cylindrical compacts of iron to permit some conclusions. Briefly, in a compact 3/4 in. high and 1/2 in. in diameter pressed at 48 tons per sq. in., the residual stresses on the surface measured by X-ray diffraction techniques are as follows: The radial stresses on top center and bottom of compacts are practically zero; the longitudinal residual stresses on the side walls are near the top quarter of the compact about seven tenths that of the yield point of mild steel in tension; at the center surface of the compact they are in compression and in magnitude about 0.3 that of mild steel in tension; near the bottom, they are practically zero. F. V. LENEL.?- I was particularly interested in the work that Messrs. Wulff and Seelig reported on the influence of lubrication on the density of parts. In connection with the work of the American Society for Testing Materials on determining the green density of a compact, we ran into considerable difficulty in getting reproducible results of green density determinations. It was much easier to get a reproducible green density when the lubricant was mixed with the powder than when the lubricant was applied to the die walls; particularly when the die was not perfect-i.e., if it was somewhat rough or out of round. In this case, it was almost impossible to get reproducible green densities when the die walls were lubricated, while it was not too difficult to get reproducible

Jan 1, 1946

Jan 1, 1970

By A. M. Gaudin

This year the American Institute of Mining, Metallurgical and Petroleum Engineers has chosen to give me its Robert H. Richards Award, a distinction which is widely regarded as the highest honor in the field of mineral engineering. I am deeply moved by this tribute, which suggests not only regard for my service to the profession but also the affectionate esteem of my fellow mineral engineers. You have made me happy at the prospect of the permanent association of my name with that of Robert Hallowell Richards and of the distinguished recipients of the award, Arthur Fay Taggart, Frank R. Milliken, John F. Myers, Edward William Englemann, C. Harry Benedict, Edward W. Davis and Arthur W. Fahrenwald. I wish to thank you all for an honor so generously bestowed. Professor Richards is known to us all as a legendary figure, and we can read about him in the Richards Memorial Volume of our Transactions. Some of us have even had the good fortune of knowing him. I count myself among those lucky few. In its intangible attributes the Award recalls those things for which Professor Richards stood: for Science in Art, for thoroughness, for meticulous accuracy, for bold imaginativeness, most of all for professional integrity. Now, as in his day, these attributes are central to an engineer's creed. They are appropriately needed in a formulation of a definition of the mineral engineering profession.

Jan 6, 1957

By J. F Haseman, J. E. Davenport

In the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1,2 results in inadequate beneficiation. Investigation of Flotation Methods Flotation has been used in 'Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate. 3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P2O5 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P2O5, 20; SiO2, 40; CaO, 29; Fe2O3, 2; Al2O3, 3; F, 2.3. The particle-

Jan 1, 1948

By J. F. Haseman, J. E. Davenport

In the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1,2 results in inadequate beneficiation. Investigation of Flotation Methods Flotation has been used in 'Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate. 3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P2O5 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P2O5, 20; SiO2, 40; CaO, 29; Fe2O3, 2; Al2O3, 3; F, 2.3. The particle-

Jan 1, 1948

By J. F. Haseman, J. E. Davenport

IN the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1-2 results in inadequate beneficiation. INVESTIGATION OF FLOTATION METHODS Flotation has been used in Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate.3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over-all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P205 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P205, 20; Si02, 40; CaO, 29; Fe203, 2; A1203, 3; F, 2.3. The particle-

Jan 1, 1947

By R. M. Kuralt

CONCESSION from the Chilean government granting the company use of the Rio Salado water stipulates that a minimum of 35,000 metric tons of such salty water must be diverted from the Salado daily, and be disposed of in a manner not to contaminate the Rio Loa to the east of its confluence with the Rio San Salvador at Chacance, Fig. 1. Consequently, the selection of a suitable site for the sulphide plant tailing disposal had the additional problem of the disposal of excess brackish waters. Studies of the terrain in the area directly to the south of the plant indicated that construction of tailing ponds would be costly, and generally unsatisfactory due to excessive slope of the ground, lack of flanking hills, danger of raising the groundwater level thus causing springs along the Antofagasta-Bolivia railroad; and the possibility of creating a dust nuisance at the plant site or at Calama due to blowing sands. Further to the southwest the San Salvador river canyon offered some possibilities and surveys were made at sites A to D, inclusive, Fig. 1. Here again costly dams would have been necessary to impound the tremendous quantity of tailing produced by the new plant, and provision would have to be made to handle a considerable flow of water through the disposal area to the west of the Rio Opache. Additional surveys were made in the area to the southeast of the plant which brought to light the existence of a natural basin of about 14 1/2 square miles, flanked by low-lying hills, and requiring only the most rudimentary dams to close the outlets at the western end. It has been estimated that the Salar de Talabre, as this basin is called, has the ultimate storage capacity of 1 billion tons of tailing with dam heights of only 60 ft. Pipe Line Layout After the decision had been reached to use the Salar de Talabre for the main disposal area various methods were considered for transporting the tailing from the 300 ft diam thickeners to the disposal site, 17.7 miles away. The first plan was to construct a. concrete launder on a 11 pct grade but a survey of this grade line indicated it would not lead directly into the Salar but would terminate somewhat to the north, and from that point on the slope would be considerably less than 1 pct along a natural watercourse. A slope of 2 pct or even 2 ¼ pct in such a rough watercourse or ditch is the ideal, judging from the success of operations at Cananea. Since a simple ditch is the most economical vehicle for the transportation of tailing, it was decided to use one with a 2 ¼ pct grade for as much of the distance as was possible with the existing terrain and to construct a concrete headbox, or forebay, at the terminus of the ditch to supply a steel pipe line leading the rest of the way into the Salar. Referring to the maps of the Chuquicamata area outlining the tailing disposal system, Figs. 1 and 2, one can trace the flow of the tailing from the thickeners to the Salar as follows: The 8-in. spigot pipes from the thickeners (1) discharge into a concrete sump feeding a concrete launder with a 2 pct slope. The launder carries the tailing into a concrete junction box (2) located south of the sulphide plant. The purpose of the junction box is to provide emergency facilities for diverting the tailing flow from the main disposal area to the emergency area. It also provides a junction point for the launder from the sulphide plant's future east extension. The box is fitted with a pair of radial gates controlling flow into the main tailing ditch (6) or the emergency ditch (3) which terminates on the pampa. All ditches are of the same cross section, namely 4 ft wide at the bottom with 1 : 1 side slopes and with a minimum depth of 21/2 ft. Where the nature of the ground demands it, the ditches are lined with stone rubble. The overflow from the plant's 10 million gal reservoir (4) is led into the main tailing ditch by way of a spillway ditch (5). The main ditch terminates in a trashscreen fitted, concrete forebay (7) which serves as a headbox for a 24 in. O.D. x 3/8 in. wall, Dresser coupled, steel pipe (10) and a 12 in. x 8-gage spiral welded, wedge-lock coupled, steel pipe (8). The forebay is fitted with two compartments, one for each pipe, and flow into each is controlled by radial gates. Each compartment is fitted with a 6-in. drain pipe controlled by a rubber pinch valve. To protect the forebay from the abrasive effects of the tailing, the sides of the compartments will be lined with old

Jan 1, 1952

By James R. Finlay

SOMETIMES the American Institute of Mining and Metallurgical Engineers appears to be a strictly technical society, and if so my paper should deal with the technical operations of finding and producing gold. But I have not been engaged in practical gold mining since 1910. What I have been doing mainly for the last 25 years is to try to figure out what the income of a mine-gold, or any other-would be worth to an investor. So I have been in contact with gold mining, in one way or another, for some 42 years. I say that with a purpose. In that time the industry has been through one of its cycles. The value of gold mines is very dif¬ferent at different points in one of these cycles. For this reason the cycle of gold mining is a highly practical study. The cycle of gold is a more or less rhythmical change in its value or purchasing power measured in the grand average of other commodities, which is the same thing as its price. Taking the price of gold in 1897 as a stand¬ard, a dollar at that time declined to 55c. in 1913, to 26.8c. in 1920, and since then it has risen to 47.4c. in 1930 and to about 64c. in 1932. It is worth about 21/ times as much now as it was in 1920, but less than two¬thirds as much as in 1897. When this change is applied to the value of a gold mine still greater differences are evident. The cost of producing gold is of course measured in labor and other commodities, and as that cost goes down the margin of profit measured in the gold itself increases. Thus a mine which might make only 10 per cent profit in 1920 on its gross output may easily make 25 per cent now. The stockholder gets 2 ½ times more gold that is worth 2 1/2 times as much in purchasing power, therefore the real value of his holding has risen by the one factor multiplied by the other, giving him 61/4 times as much goods or services as he had in 1920.

Jan 1, 1933

By P. S. Bagga, P. T. Luckie

Liberation (the process of destroying the interlock between unwanted materials, such as mineral matter and pyrite, and coal) is one of the most important precursors to the benefication of raw coal in order to produce high grade marketable coals at economical recovery levels. Liberation is achieved usually by comminuting the raw coal. However, because of the significant interaction between the production of fine particles to achieve liberation and the cost of separating fine particles, the coal industry needs a method projecting the change in washability with size reduction. Such a projection can be combined with size reduction and separation simulators in order to examine completely beneficiation circuitry to produce coals of stringent specification. This paper presents the development of an appropriate liberation model for coal and its solution by Monte Carlo simulation to predict the expected washability of the raw coal as a function of comminution. It includes model validation with a known binary mixture as well as with a known teritary mixture.

Jan 1, 1983

Jan 1, 1967

CYRIL STANLEY SMITH, Chairman CARL E. SWARTZ, Past-Chairman ARTHUR PHILLIPS, Vice-Chairman E. E. SCHUMACHER, Vice-Chairman H. A. MALONEY, Treasurer FRANK. T. SISCO, Secretary 9 West 39th Street, New York, N. 1. Executive Committee 1944 1945 WILLIAM C. ELLIS C. S. BARRETT ALAN MORRIS T. A. WRIGHT KENT R. VAN HORN L. L. WYMAN

Jan 1, 1943

For the benefit of those of our members who are considerable holders of securities,but owing to their isolated situations are not in close touch with the metropolitan market and current quotations, we are publishing two very interesting graphs, showing the course of the bond and stock markets in New York over a long period. These graphs are furnished through the courtesy of the financial department of the New York Tribune.

Jan 11, 1919

LIST OF MEMBERS AND ASSOCIATES ARRANGED ACCORDING TO STATES AND TOWNS. ALABAMA. Aldrich, Aldrich, W. F. Anniston, Noble, A. E Bessemer, Fergusson, V. Bir¬mingham, Aldrich, T. H., Jr, Bowron, C. E, Bowron, J., Davis, F H, Dimmick, F. D., Hillman, Ludwig, C. P., McCarthy, M. E, McCormack, G B., McCrery, Ramsay, E, Shook, A. M. Ensley, McDonald, S., Shook, J. W. Gadsden, Lee, R. H., Jr. Gannts Quarry, Hiller, E Helena, Squire, J. Huntsvalle, Moss Le¬high, Toulmin. Ozmoor, Greenamyer, A. G. Rock Run, Garvin, J M. Selma, Saylor, B. F. A. Sheffield, Miller, D, I. Tuscaloosa, Lodge, J. Woodward, Wood¬ward, A. H. ALASKA Coppermount, Wiswell, H. J. Douglas Island, Macdonald, J Ellamar, Stevens, B. Fairbanks, Hamilton, A L Haines, Le Wald, E A. Jualin, Hoggatt, W. B. Juneau, Denby, Hampton, W. H , Thane, B L , Whipple, J. R. Ketchikan, Bar¬ber, J C , Parker, J. L., Reid, G D Nome, Garfield, C D. Snettwham, Tisdale. Teller, Hirshberg, M. R. Treadwell Caty, Kinzie, Stow, H. P. Valdez, Esterly, G. M, Hazelet.

Jan 1, 1910

Mining Methods JAY A. CARPENTER, Chairman CLYDE E. WEED, Vice-Chairman ROY A. HARDY, Secretary Alluvial Mining CLINTON BERNARD O. B. PERRY NORMAN CLEAVELAND C. M. ROMANOWITZ R. B. EARLING J. B. STAPLER BENERE H. GRANT CHARLES THURMAN Open-Cut Mining Iron G. E. DIEHL EARL E. HUNNER P. G. HARRISON G. E. LEVEQUE Copper LOUIS BUCHMAN W. H. GOODRICH J. H. DAVIS WALTER S. LARSH

Jan 1, 1941

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

ENGINEERS AVAILABLE (Under this heading will be published notes sent to the Secretary of the Institute by members or other persons introduced by members.) Member, technical graduate, aged 36. Experienced copper metallurgist and mining geologist. Familiar also, through 15 years' broad practical experience, with mining, cyanidation, concentration (including oil flotation), metallurgical accounting, and mine and plant management. Best references. No. 237. Member, graduate engineer with 10 years' experience in malleable, gray -Iron, and steel, as chemist., metallographist, and metallurgist. Researches and investigations by laboratory methods and under works conditions. No objection to some routine. No. 243. Member, technical graduate, broad experience in coal, oil, and clay. Will go anywhere as mining geologist after Sept. 1, 1915. No. 244. Member, technical graduate, aged 30, married, with over 10 years' experience as chemist, engineer on construction, superintendent and general superintendent of bituminous coal and coke plants. SpeaksSpanish. References. No. 245. Member, open for engagement as mine safety inspector or as superintendent of bituminous coal mine. No. 246. Member, graduate mining engineer, with five years' experience in coal mining as mining engineer, construction engineer, pit boss, and superintendent, open for engagement. Best references. No. 247.

Jan 9, 1915