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By Walter S. Lutz, J. T. Griffith

The rainfall during the last six months of 1942 in Scranton and vicinity was 24.06 in.—only 4.35 in more than the average for this district in any equivalent period—yet that rain forced The Hudson Coal Co. to handle 38.8 tons of water for every ton of coal mined. It filled to overflowing, or to a higher level, 14 flooded mine areas holding, in total volume, 8.3 billion gallons of water in storage; brought about a pumpage of 11.4 billion gallons in the months of December 1942 and January, February and March 1943 Half of the total energy generated or purchased (44,322,980 kw-hr.) was required to keep the mines sufficiently free of water to permit the mining of coal during a period of much needed output. Fig. I shows the extent to which this flooding of exhausted and idle properties has proceeded, and is known to be proceeding further, at the upper end of the Northern field. This rainfall (24.06 in.) means 80,559,444 tons of water on approximately 46.2 sq. miles of area tributary to the pumping plants. During the high rainfall period, 72,687,661 tons went underground, being pumped at the stations in this territory. This means an inflow of 88per cent, a runoff of 12 per cent compared with 64 per cent during normal conditions in a normal region. During this period no floods, either surface or underground, had to be dealt with, although the streams and rivers were high on several occasions. In general, the rainfall for the first three months of 1943 can be described as heavy but not abnormal.9 These figures not only indicate the large volumes of water to be handled but, by their very size, they also emphasize that the mine-water problem is not to be conquered, at least from an economical standpoint, by merely buying and installing more and more pumps. Into the solving of this water problem must go much planning, adhering to the principle of keeping the highest percentage of water at the highest possible level for the longest period of time; the storage of huge volumes in sumps capable of being reduced after the flood inflow and at rates less than the inflow; plans for prevention of surface inflow; provision of the minimum of pumping capacity arranged to as nearly 100 per cent pumping load factor as is possible; intercompany joint pumping plants; the conducting of water from one plant to another and from one property to another over distances measured in miles; priorities in power to the threatening situations; and. balance of the necessities and dangers existent in one colliery over another, day by day, when quantities of water are running high. Despite these astronomical figures of water volumes, their disposal to the surface is accomplished reasonably cheaply because, although the annual cost of $575,020 is a large sum, the cost per ton is about 0.1168 cents. Credit for this goes

Jan 1, 1947

By J. T. Griffith, Walter S. Lutz

The rainfall during the last six months of 1942 in Scranton and vicinity was 24.06 in.—only 4.35 in more than the average for this district in any equivalent period—yet that rain forced The Hudson Coal Co. to handle 38.8 tons of water for every ton of coal mined. It filled to overflowing, or to a higher level, 14 flooded mine areas holding, in total volume, 8.3 billion gallons of water in storage; brought about a pumpage of 11.4 billion gallons in the months of December 1942 and January, February and March 1943 Half of the total energy generated or purchased (44,322,980 kw-hr.) was required to keep the mines sufficiently free of water to permit the mining of coal during a period of much needed output. Fig. I shows the extent to which this flooding of exhausted and idle properties has proceeded, and is known to be proceeding further, at the upper end of the Northern field. This rainfall (24.06 in.) means 80,559,444 tons of water on approximately 46.2 sq. miles of area tributary to the pumping plants. During the high rainfall period, 72,687,661 tons went underground, being pumped at the stations in this territory. This means an inflow of 88per cent, a runoff of 12 per cent compared with 64 per cent during normal conditions in a normal region. During this period no floods, either surface or underground, had to be dealt with, although the streams and rivers were high on several occasions. In general, the rainfall for the first three months of 1943 can be described as heavy but not abnormal.9 These figures not only indicate the large volumes of water to be handled but, by their very size, they also emphasize that the mine-water problem is not to be conquered, at least from an economical standpoint, by merely buying and installing more and more pumps. Into the solving of this water problem must go much planning, adhering to the principle of keeping the highest percentage of water at the highest possible level for the longest period of time; the storage of huge volumes in sumps capable of being reduced after the flood inflow and at rates less than the inflow; plans for prevention of surface inflow; provision of the minimum of pumping capacity arranged to as nearly 100 per cent pumping load factor as is possible; intercompany joint pumping plants; the conducting of water from one plant to another and from one property to another over distances measured in miles; priorities in power to the threatening situations; and. balance of the necessities and dangers existent in one colliery over another, day by day, when quantities of water are running high. Despite these astronomical figures of water volumes, their disposal to the surface is accomplished reasonably cheaply because, although the annual cost of $575,020 is a large sum, the cost per ton is about 0.1168 cents. Credit for this goes

Jan 1, 1947

By F. M. Smith, R. H. Moore, J. R. Morrey

Electrodiffusion in y cerium reported by Henrie has been confirmed and a Preliminary estimate made of the relative rates of electrodiffusion of iron, cobalt, and nickel. These diffuse to the anode at rates decreasing in that order. In addition, copper and manganese exhibit slow, but detectable, diffusion to the anode and molybdenum exhibits detectable diffusion to the cathode. The electrodiffusion of carbon, .zirconium, antimony, magnesium , and silicon in y cerium could not be detected. Iron and cobalt diffuse in y cerium at rates proportional to the current density and with no apparent dependence on temperature. Decreasing polarization of iron and cobalt with increasing temperature, which cancels the expected rate increase, would account for this behavior. The electrodiffusion rate of iron in y uranium and in E plutonium has been measured. Diffusion of iron is anode-directed. Tin was found to diffuse to the cathode, in y uranium, at an appreciable rate. In all of these solvent metals, negative ions diffuse to the anode and positive ions to the cathode. The potential field effect appears to account satisfactorily for these results. FROM early experimental work summarized by Jost1 and Seith,2 the driving force for electrodiffusion was attributed to the potential field acting on ions in a metal. More recently, Heuman,3 Wever,4 and Huntington5 have shown that momentum interchange between conduction electrons and mobile entities in the metal contributes to electrodiffusion. Electron momentum interchange is anodically directed and the direction of diffusion resulting from the field force is dependent upon the charge on the diffusing entity. These two effects may either reinforce or oppose each other. Glinchuk6 has pointed out that momentum transfer in defect conductors should be cathode-directed and this appears to be the case as demonstrated by wever's4 work on iron. Barnett's7 work, on the other hand, indicates that, even in defect conductors, electrons show a negative E/m ratio when accelerated with respect to the lattice and should lead to anode-directed momentum transfer. In discussing this problem, Wever and seith8 suggest that defect electrons interact preferentially with activated ions so as to allow a net movement toward the cathode while still maintaining an electron momentum transfer in the anode direction. Williams and Huffine9 and Henriel0 have demonstrated that electrodiffusion may be useful for purification of yttrium and cerium. In yttrium, Williams and Huffine note that movement of several metallic impurities toward the anode is in keeping with observations in most other metallic systems and indicates that yttrium remains a normal electronic conductor at least to 1230°C. Close inspection of their data shows, however, that oxygen, nitrogen, and the transition elements diffused toward the anode, while nontransition elements diffused toward the cathode. This suggests that potential field effects may have been appreciable. The present work was concerned with the applicability of electrodiffusion as a technique for purification of plutonium, but, because of the obvious hazard inherent in work with this metal, experiments to develop the technique were carried out using cerium and uranium. The results of electrodiffusion measurements on these metals and on plutonium are reported here. EXPERIMENTAL The metal specimens prepared for this work were 6 in. long, 1/4 to 1/2 in. wide, and 1/16 to 3/32 in. thick. The uranium specimens were machined from a bar which analyzed 310 ppm Fe and the electrodiffusion of iron was followed by spectrographic and by chemical analysis. Cerium and plutonium specimens were cut from sheet rolled from ingots obtained from molten salt-metal equilibrations during which radioactive tracers were introduced. The electrodiffusion of the tracers was subsequently determined by counting methods. The specimens were electrolyzed between nickel electrodes containing resistance heaters used to equalize the specimen and electrode temperatures, thereby reducing thermal gradients. The temperature of the electrodes adjacent to the ends of the specimen was measured with chromel-alumel thermocouples which were connected to the heater controls. The surface temperature of the specimen at a point midway between the electrodes was measured with a sapphire rod pyrometer, the output of which controlled the dc power supply. This assembly was enclosed within an evacuable chamber containing a quartz viewing window. The temperature of the specimen over its entire length could be scanned with a portable pyrometer through

Jan 1, 1965

By R. R. Vandervoort, W. L. Barmore

Tensile creep experiments were conducted on high-purity, poly cvystalline rhenium from 1500" to 2300°C at stresses from 1500 to I0,OOO psi in a vacuum of 10-a torr. The apparent activation energy for creep was 60 kcal per mole, and the steady-state creep rate varied directly with stress to the 3.4 power. Dislocation substructure that developed during creep was studied by transmission electron microscopy. Possible rate-controlling deformation mechanisms are discussed. The creep behavior of most metals at elevated temperature can be represented by the following equation:&apos;&apos;&apos; t = Cf(s)(^)(s/E)nD [1] where i = steady-state creep rate, C = constant, f(s) = a function involving microstructure, s = applied stress, E = the average elastic modulus at test temperature, n = constant, D = diffusion coefficient According to this well-established relationship, metals with higher elastic moduli and lower diffusion coefficients should have greater creep resistance at the same stress and temperature and equivalent mi-crostructures. While no diffusion data are available, the diffusivity of rhenium should be less than that for most other refractory metals because of its high melting point and hcp crystal structure. The Sherby-Simnad relation for calculating atomic mobility in metallic systems3 predicts that the diffusion coefficient for rhenium is less than that experimentally determined for tungsten4 in the temperature region 1500. to 2200°C. At these temperatures the elastic modulus for tungsten5 is only slightly larger than the extrapolated modulus for rhenium.6 Thus, rhenium is a good possibility for a a high-temperature structural material, but few data on the creep of rhenium have been reported. This investigation was undertaken to study the high-tempera-ture deformation behavior of rhenium in detail. EXPERIMENTAL TECHNIQUES The material used in this study was consolidated from high-purity powder. After cold pressing the powder to a plate a in. thick, the billet was sintered in hydrogen at 2250°C for 24 hr. The plate was reduced to 0.100 in. by cold cross rolling with intermediate anneals at 1650°C for 20 min between passes. The plate was further reduced to 0.060 in. by unidirectional cold rolling with similar heat treatments between passes, and then finally stress-relieved in hydrogen at 1650°C for 30 min. Specimens tested at 1900°C and below were pretest-annealed at 1900°C for 2 50 hr. Specimens tested above 1900°C were pretest-annealed at 2400°C for 5 hr. The impurity content in the "as-received" plate was quite low, table I. Essentially no change in impurity levels was detected in specimens after creep testing. All creep tests and annealing treatments were conducted in a vacuum of 10-8 torr in a test furnace heated by a tungsten mesh element. The load was applied to the specimens through a bellows, and stresses were maintained to ±1 pct of the selected value by periodic corrections for changes in specimen cross-sectional area during creep and for changes in the bellows spring force due to load column extension. One-inch-diameter tungsten force rods were used in the hot zone of the furnace. Deformation at temperature was measured by optically tracking gage marks on the specimen. Temperature was measured by a calibrated optical pyrometer and was determined to ±5"C. Grain sizes were determined by the linear intercept method and specimens were examined in the "as-polished" condition, using polarized light. Specimens annealed at 1900°C had a grain size of 52 ± 5µ , and those annealed at 2400°C had a grain size of 148 * 11 µ. Pieces were cut from the gage section of creep-tested specimens and planed to a thickness of about 0.010 in. by spark discharge machining. Thin foils for viewing by transmission electron microscopy were obtained by electropolishing in a solution of 6:3:1 ethyl alcohol, perchloric acid, and butoxy ethanol, respectively, using the window technique. Bath temperature was —4OoC, and the cell potential was 35 v. The foils were examined in Siemens Elmiskop I, operating at 100 kv. RESULTS AND DISCUSSION In order to analyze the results from creep experiments, Eq. [I] is rewritten in the following form: <=Kf(s)ne-/RT [2] where K = constant, ?// = apparent activation energy for creep,

Jan 1, 1970

By Nicholas J. Grant, A. W. Mullendore, Paul Predecki

The Duwez technique of splat cooling in which a molten droplet of metal is accelerated and made to impact on a cold, highly conducting substrate was investigated.- An apparatus for producing "splat" was constructed employing an explosive powder charge to accelerate the molten metal. Transport of the molten metal just prior to impact with the substrate was studied by means of high-speed photography. The molten particles are small spherical droplets from about 1 to 50 µ diameter. The average cooling rates for aluminum, silver, and a gold alloy splatted on nickel substrates at room temperature were determined experimentally and were found to vary from 18 to 5 x 10&apos; C per sec. The heat-transfer coeficients for pure aluminum and pure silver splats cooled on nickel substrates at room temperature were found to be 2.7 to 6.8 and 13.6 to 54.2 cal per sec sq cm &apos;C (2 to 5 x 104 and 1 to 4 x 18 Btu per hr sq ft OF), respectively. Solidification rates in pure aluminum and silver splats were calculated. "SPLAT cooling" is a term describing a technique for extremely rapid freezing and cooling of molten metals and alloys to room temperature or below. The technique was developed by Duwez et al.&apos; in 1960, and after refinement2 consisted in transferring a few tenths of a gram of molten metal to near sonic velocity to strike a suitably placed cold copper surface. Upon impact, the metal spread into a thin nonuniform film called a splat, about 10-4 cm thick. The splat particles produced in this manner were thin enough in some areas to be suitable for transmission electron microscopy, without further treatment, and, together with X-ray data, yielded a variety of structures which would be classified as follows: 1) supersaturated solid solutions (increase in solubility limit), 2) metastable crystalline stoichiometric and non-stoichiometric intermediate phases ( in Au-20.5 at. pct Si and in Au-14 at. pct Sb, respectively), 3) amorphous alloys, 4) retained high-temperature phases, 5) alloys having equilibrium phases present, but with unusual, markedly altered structures. The crystalline phases present in splats were usually extremely fine-grained and had low dislocation densities. Present interest in splat cooling centers around a study of the unusual structures produced by the technique, their contribution to alloy theory, and the possibility of the development of new or unusual properties. In addition, the technique is being examined for the production of bulk quantities of new high-strength alloys for low- and high-temperature use. Although a number of alloy systems have already been investigated by the technique, relatively little is known about the mechanism of splat formation and the physical conditions encountered by the metal during solidification and cooling. The purpose of this work was: 1) to determine the velocity, shape, and size of the molten metal droplets just prior to impact with the cold substrate; 2) to try to estimate or measure solidification rates, heat-transfer coefficients, and cooling rates in splats. EXPERIMENTAL TECHNIQUE Splatting Apparatus and Procedure. An apparatus similar to the type originated by Duwez et a1.,2 but employing an explosive charge to accelerate the molten metal, was used. A "Ramset" fastening tool (a gun normally used for driving studs into concrete, steel, and so forth) was mounted vertically above a resistance-heated graphite crucible as shown schematically in Fig. 1. A splat product was produced by melting a few tenths of a gram of metal at the position shown in Fig. 1, and then exploding a powder charge in the breach. The shock wave thus generated traveled down the barrel into the furnace, ejecting the molten metal through the 0.06-in.-diameter hole at the bottom of the crucible. The ejected metal was impacted on a high-conductivity metallic substrate where it formed a splat. High-Speed silhouette Photography. In order to investigate the size and shape of the ejected, molten metal just prior to impact with the cold substrate, high-speed silhouette photographs were taken, and are shown in Figs. 2 and 3. These photographs were obtained with the aid of an Edger-ton, Germeshausen, and Grier microflash unit and a submicrosecond flash drive equipped with a variable time delay. Two types of triggers were

Jan 1, 1965

By W. F. Hower, W. Brown

Large-scale sand consolidation tests were conducted in an effort to determine the reasons for the successes and failures of this method of sand control. Several different consolidating materials were used in treating both clean and bentonitic sands that were packed in a chamber having a capacity of 3.3 cu ft. The results were essentially the same for all of the different consolidating materials, The data show that low-viscosity consolidating materials pumped at a relatively slow rate gave the best results. Where the formation has produced sand, the treating fluids can compress the formation, thus permitting the channeling of fluids to another horizon. Pressure-packing these zones before attempting to consolidate is recommended. Sands containing more than 4 per cent of water-swelling clays are not good candidates for consolidation. It is indicated that loose sand, particularly when it is bentonitic, can be fractured during the placement of the treating fluids. INTRODUCTION Sand production in oil and gas wells has plagued the industry for many years, and numerous cures for this problem have been suggested. Most methods have been successful to a certain degree, but the great variety of well conditions that exist in the different areas has magni- fied the problem and limited the successful use of the various systems. Four review papers1-4 present a wealth of information concerning the degrees of success that have been obtained by the different sand-control methods. The bridging of sand grains by the use of gravel packs and screens has been quite successful. However, these methods do not leave the casing clear for all types of multiple completions, and the cure does not last for the production life of the well in some instance:;. The control of loose sands by sand consolidation with resins has never been as successful as desired. It has always been hoped that such a treatment would eliminate all sand problems for the life of the well, but. initial applications, starting in the middle 1940&apos;s, were only moderately successful. Lott, et a1,3 reported a success ratio of approximately 50 per cent and made the following conclusions. The highest percentage of successes were obtained where: a. Consolidation of a zone was made at the time of initial completion or prior to the production of sand. b. The interval treated was less than 12 ft in length. c. Between 30 and 50 gill plastic/ft of producing interval was displaced through the perforations. REASONS FOR SAND CONSOLIDATION FAILURES Our own experiences in the field of sand consolidation point toward the following conditions as the major reasons for the failure of sand consolidation attempts. 1. Mud-plugged perforations and mud invasion of the formation. 2. Sand in the casing covering all or part of the perforations. This sand could be either formation sand or one of the coarser sands used as propping agents in hydraulic fracturing. 3. Holes in the casing. 4. Channels behind the casing. 5. Attempting to treat too long a perforated section. 6. Too high a percentage of water-swelling clays in the formation. 7. Formations that have produced sand. Recent attempts were made to treat perforated sections ranging from 10 to 30 ft, in wells that have produced sand, by using a straddle packer that was raised and lowered through the perforations as the consolidating material was being pumped. In most instances, the pressure required to pump fluid into the formation varied considerably as the tool was raised and lowered. This suggested the possibility that significant differences in permeability were present or that only part of the formation had produced sand. There were times when a sudden break in pressure indicated that a fracture was being formed. Research conducted several years ago concerning the problem of the control of water in air and gas drilling indicated that shale sections could be fractured quite easily. In addition, it was determined that it was easier to pump fluids into shale bodies by fracturing the shale itself, or the interface between the shale and sand, than to pump into a fluid-saturated formation. Formations that produce sand are usually adjacent to shale bodies and frequently have shale streaks of various thicknesses inter-bedded in the sand. Therefore, where shale is exposed to fluid pressure it

By A. Stanley, J. C. Mead

The MacIntyre Development of the National Lead Co. is located at Tahawus, N. Y. The operations involve the mining and concentrating of a titaniferous iron ore to produce an ilmenite concentrate and a magnetite concentrate. Construction of the MacIntyre plant was commenced during the summer of 1941,when world conditions threatened to cut off the supply of Indian ilmenite. An open pit mining operation was developed and the crushing and milling equipment put in operation in July 1942. A general description of the operation was given in the Adirondack Issue of Mining and Metallurgy for November 1943. The metallurgy of the mill operation was described by Mr. Frank R. Milliken,* Plant Manager, National Lead Company, MacIntyre Development, and presented at the AIME New York Meeting, February 1948. The magnetite concentrate produced in the milling operation was too fine (minus 20 mesh) to be used directly in iron blast furnace operation, and most of the magnetite had to be stockpiled in 1942 and 1943. In 1943, the Defense Plant Corp. built a Greenawalt sintering plant at Tahawus, N. Y., to put the magnetite concentrate in a more suitable form for use in the iron blast furnace. The Greenawalt sintering plant consists of three 10 by 25 ft sintering pans designed to produce 1800 gross tons of sinter per 24 hr. The vacuum to each pan is produced by two Greenawalt fans in series, pulling approximately 30,000 cu ft of air per minute at 50 in. water gauge vacuum. The plant started operation in August 1944. The present plant production averages 25 tons per operating pan hour (approximately 224 lb per operating hour per square foot of grate area) of plus 1 in. sinter. Raw feed to the plant consists of 61 pct magnetite, 4 pct anthracite coal culm, and 35 pct minus 1 in. return fines which are conveyed to a pug mill where the materials are mixed thoroughly and water added to give the mixture 5.5 to 6 pct moisture. The mixed prepared feed is conveyed to two 4 by 10 ft vibrating screens where the minus 1 in. plus 5/8 in. return fines are screened out and discharged into a surge bin for use as a hearth layer. The minus % in. prepared feed is discharged into another surge bin for use as prepared feed. A charge car, electrically operated, having a capacity of one charge of prepared feed and several charges of hearth layer, lays a thin layer of plus 5/8 in. return fines and 9 1/2 in. depth of prepared feed into the pans. A fluffing roll and a vibrator on the car fluffs and spreads the prepared feed into the pans. An ignition car, electrically operated, ignites the top of the bed with a 30 sec flash burn. The 9 1/2 in. bed sinters in approximately 13 min. Dumping the pan, and recharging and igniting the bed requires 2 min. To improve the quality of the ilmenite concentrate produced in the mill and to reduce the amount of titanium dioxide lost in the mill tailings and in the magnetite product, extensive research work and pilot plant operations have been done on grinding the crude ore to minus 65 mesh size (rather than to minus 20 mesh) and concentrating it by a combination of magnetic separation (for magnetite recovery) and flotation (for ilmenite recovery). These tests have proved successful in increasing ilmenite recovery and grade. With the development of the ilmenite flotation process to a stage where a full scale flotation plant was in the design stage, the problem arose of handling the 65 mesh magnetite concentrate that would be produced. In order to study and solve the problems of handling and sintering the 65 mesh magnetite in the sinter plant, a pilot sinter plant was plus from John E. Greenawalt. The effect of using 65 mesh magnetite in the sintering operations was then studied on the 2.4 sq ft test pan, operating under conditions as similar to the large plant as could be set up in the laboratory. A series of tests were run in the test pan on present sinter plant feed that had been mixed in the plant pug mill. An average production and an average quality of sinter produced in this series

Jan 1, 1950

By H. C. Casey, M. B. Panish

The roles of the phase diagram and the diffusion and solubility analyses in the selection of sources for the diffusion of zinc into GaAs are discussed. Isothermal sections of- the phase diagram are described at 1000°, 775°, 744°, and 700°. Below 744° the three solid phases, Zn3As2, GaAs, and ZnAsz, may be in equilibrium. In the composition regions where the three solid phases are in equilibrium, the partial pressures of the components do not vary with composition and there is no liquid phase. A diffusion source composition within that region containing 5, 50, and 45 at. pct Ga, As, and Zn was selected to yield reproducible diffusion profiles without surface damage. The 5/50/45 ternary source Provides a significant improvement in the reproducibility and planarity of- the zinc diffusion as compared with other zinc diffusion sources that result from starting with elemental zinc, dilute solutions of zinc in gallium, or various combinations of zinc and arsenic. Several experimental diffision profiles for the 5/50/45 diffusion source have been determined. The profiles are very steep in the region of the junction and produce devices that exhibit step junction behavior. The junction depth varies as the square root of time. Diffusion at 700°C for 4 hr gives a 5-µ junction depth and a surface concentration of 2.4 x1020 cm-3. The surface concentration is enhanced by a factor of three and the junction depth decreased by a factor of ten as compared with diffusion with a starting source 01- elemental zinc. The differences in dijyusion behavior with these two sources are due primarily to the approximately 10&apos; greater equilibrium As4 partial pressure for the 5/50/45 source. ThE most common technique for the preparation of GaAs p-n junctions for diverse applications ranging from switching diodes to injection lasers has been to diffuse zinc into the solid from the vapor phase. The zinc impurity has generally been provided by introducing into a diffusion ampoule elemental zinc, dilute solutions of zinc in gallium, or various combinations of zinc and arsenic. For the purpose of eliminating surface deterioration, zinc has been diffused into GaAs through SiO2 films1 and from zinc-doped SiO2 films.2 During a study of the injection and recombination mechanisms in GaAs junctions, we found that these diffusion techniques either resulted in nonreproducible diffusion profiles and nonplanar junction interfaces or were extremely inconvenient to use. An unpublished report by Allen and pearson3 emphasized that an understanding of diffusion requires relating the diffusion conditions and system compositions to the ternary phase diagram. Recent studies have made it possible to relate the effect of starting source composition on the GaAs sample surface and diffusion profile reproducibility and, in addition, to estimate the expected surface concentration and zinc diffusivity. These studies include the Ga-As-Zn phase diagram,415 the determination of the incorporation reaction for zinc as a substitutional impurity,6 and the analysis of the concentration dependence of the diffusion coefficient.&apos; In this paper the role of the phase diagram in the selection of the starting diffusion source composition is illustrated, and an application of the diffusion and solubility analyses is used to demonstrate their utility in the estimation of surface concentration and diffusivity for various starting source compositions. Consideration of the isothermal sections of the ternary phase diagram shows that below about 744° C a composition region exists where the three solid phases Zn3As2, GaAs, and ZnAs2 are in equilibrium without a liquid phase. In this region the partial pressures of the components do not vary with composition, and thus the surface concentration and diffusivity depend only on temperature. A diffusion source composition within that region containing 5, 50, and 45 at. pct Ga, As, and Zn was selected. As will be described in detail, this diffusion source yields reproducible diffusion profiles without surface damage, planar junction interfaces, and steep profiles that result in devices with step junction behavior. Only closed diffusion systems are considered. Open or flowing vapor systems require the maintaining of constant vapor pressures in the flowing gas and represent control problems outside the scope of this discussion.

Jan 1, 1969

By D. G. Westlake

Alloys of poly crystalline Nb-H have been tensile tested at 77" and 120°K after slow-cooling and after quenching from room temperature. A rationale has been developed to explain the effect of cooling rate on ductility in terms of microstructure. The inferior ductility of the slow-cooled alloys in this temperature range can be attributed to the precipitation of coarse hydride particles that fracture in the early stages of plastic deformation. Quenching Produced much smaller hydride particles and a decreased tendency for initia-tion and popagation of cracks. Prismatic dislocation loops, punched into the matrix of quenched alloys, eliminate the yield point and may Play some role in the low temperature ductility of the quenched alloys. The difference in ductility between quenched and slow-cooled V-H alloys was even greater than that observed in the N b-H alloys. It is well known that the ductility of Nb-H and V-H alloys decreases with increasing hydrogen concentra-tion and with decreasing temperature.1-l1 The follow-ing are some of the numerous mechanisms that have been adopted in attempts to explain these observations: 1) hydrogen-dislocation interarction,4&apos;6 2) lowering of the crack surface energy by adsorbed hydrogen,9&apos;10 3) localized lowering of the niobium bond strength,9 and 4) exertion of pressure inside voids by molecular hydrogen.6 A recent study1&apos; indicates that, in some way, the embrittlement is related to the precipitation of a hydrogen-rich phase, because embrittlement and hydrogen solubility have the same temperature dependence. Most of the investigators observed that at very low temperatures there was a return of the ductility. Thus, hydrogenated niobium and vanadium may be said to exhibit a low temperature ductility minimum."1&apos;8&apos;9 It has been reasoned by most observers that diffusion is required for the embrittlement mechanism to be operative, and that as hydrogen becomes immobile at very low temperatures, the ductility increases. This investigation of the mechanical behavior of Nb-H al-loys under various conditions is an attempt to gain a better understanding of the low temperature ductility minimum. EXPERIMENTAL Tensile specimens, with gage length dimensions 20 by 3 mm, were machined from sheet stock" and both faces were ground on wet 600 grit Sic paper to obtain a final thickness of 0.5 mm. Annealing as described earlier12 resulted in average grain diameters of 0.3 mm. The recrystallized niobium tensile specimens were electropolished in 90 parts H2S04, 10 parts HF (48.9 pct) by volume, at 287°K and 30 v. Hydrogen, which may have been present in the starting material and/or added during specimen preparation, was removed by annealing at 1073°K in a dynamic vacuum of 2 X 10-6 torr. Hydrogenated specimens (1.15 at. pct H) were prepared by reaction with a known quantity of pure hydrogen (obtained by thermal decomposition of UH3) at 1073oK, followed by cooling at the rate of 100" per hr. These were vacuum encapsulated, homogenized at 573°K for 3 to 5 hr, and furnace-cooled. Tensile tests were conducted on an Instron machine at a strain rate of 4 x l0-5 per sec. The gripping device was such that the specimen was positioned below the crosshead. Thus, subambient test temperatures could be obtained by immersion of the device in a cooling bath of either liquid nitrogen or Freon-12. The temperature of the Freon-12 was controlled within ± 0.1°K of the desired temperature by automatically monitoring the flow of liquid nitrogen through a copper coil in the bath. Slow-cooling to 77°K was accomplished as follows. The testing device was surrounded by a rubber jacket, open at the top and with small holes in the bottom. Liquid nitrogen, brought up to the bottom of the jacket, entered the holes and vaporized quickly. The temperature was monitored at the specimen with a copper-constantan thermocouple, and a constant cooling rate of 2 deg per min was easily achieved by controlling the level of the liquid nitrogen outside the jacket. RESULTS AND DISCUSSION Low Temperature Ductility Minimum. Resistometric studies of both Nb-H12 and V-H13 indicated that at tem-peratures less than 150°K, virtually all of the hydrogen

Jan 1, 1970

By Joseph C. Mead, Alan Stanley

The MacIntyre Development of the National Lead Co. is located at Tahawus, N. Y. The operations involve the mining and concentrating of a titaniferous iron ore to produce an ilmenite concentrate and a magnetite concentrate. Construction of the MacIntyre plant was commenced during the summer of 1941,when world conditions threatened to cut off the supply of Indian ilmenite. An open pit mining operation was developed and the crushing and milling equipment put in operation in July 1942. A general description of the operation was given in the Adirondack Issue of Mining and Metallurgy for November 1943. The metallurgy of the mill operation was described by Mr. Frank R. Milliken,* Plant Manager, National Lead Company, MacIntyre Development, and presented at the AIME New York Meeting, February 1948. The magnetite concentrate produced in the milling operation was too fine (minus 20 mesh) to be used directly in iron blast furnace operation, and most of the magnetite had to be stockpiled in 1942 and 1943. In 1943, the Defense Plant Corp. built a Greenawalt sintering plant at Tahawus, N. Y., to put the magnetite concentrate in a more suitable form for use in the iron blast furnace. The Greenawalt sintering plant consists of three 10 by 25 ft sintering pans designed to produce 1800 gross tons of sinter per 24 hr. The vacuum to each pan is produced by two Greenawalt fans in series, pulling approximately 30,000 cu ft of air per minute at 50 in. water gauge vacuum. The plant started operation in August 1944. The present plant production averages 25 tons per operating pan hour (approximately 224 lb per operating hour per square foot of grate area) of plus 1 in. sinter. Raw feed to the plant consists of 61 pct magnetite, 4 pct anthracite coal culm, and 35 pct minus 1 in. return fines which are conveyed to a pug mill where the materials are mixed thoroughly and water added to give the mixture 5.5 to 6 pct moisture. The mixed prepared feed is conveyed to two 4 by 10 ft vibrating screens where the minus 1 in. plus 5/8 in. return fines are screened out and discharged into a surge bin for use as a hearth layer. The minus 5/8 in. prepared feed is discharged into another surge bin for use as prepared feed. A charge car, electrically operated, having a capacity of one charge of prepared feed and several charges of hearth layer, lays a thin layer of plus 5/8 in. return fines and 9½ in. depth of prepared feed into the pans. A fluffing roll and a vibrator on the car fluffs and spreads the prepared feed into the pans. An ignition car, electrically operated, ignites the top of the bed with a 30 sec flash burn. The 9½ in. bed sin-ters in approximately 13 min. Dumping the pan, and recharging and igniting the bed requires 2 min. To improve the quality of the ilmenite concentrate produced in the mill and to reduce the amount of titanium dioxide lost in the mill tailings and in the magnetite product, extensive research work and pilot plant operations have been done on grinding the crude ore to minus 65 mesh size (rather than to minus 20 mesh) and concentrating it by a combination of magnetic separation (for magnetite recovery) and flotation (for ilmenite re-covery). These tests have proved successful in increasing ilmenite re-covery and grade. With the development of the ilmenite flotation process to a stage where a full scale flotation plant was in the design stage, the problem arose of handling the 65 mesh magnetite concentrate that would be produced. In order to study and solve the problems of handling and sintering the 65 mesh magnetite in the sinter plant, a pilot sinter plant was secured from John E. Greenawalt. The effect of using 65 mesh magnetite in the sintering operations was then studied on the 2.4 sq ft test pan, operating under conditions as similar to the large plant as could be set up in the laboratory. A series of tests were run in the test pan on present sinter plant feed that had been mixed in the plant pug mill. An average production and an average quality of sinter produced in this series

Jan 1, 1950

By Jack Washburn, E. R. Parker

Kinking in zinc single-crystal tension specimens was observed under conditions of low stress and high temperature. Kinking is discussed in relation to other plastic bending phenomena on the basis of dislocation theory. Experiments on the stress-induced motion of small angle boundaries are reported. KINKING was first reported by Orowan1 as a new deformation mechanism. He observed the phenomenon in cadmium single crystals loaded in compression. Glide lamellae in a local region of the rod were observed to snap over suddenly to a tilted position, resulting in a sudden shortening of the specimen. Between the tilted portion and the rest of the crystal there were fairly sharp boundaries, with the slip plane assuming approximately mirror image positions on opposite sides of these boundaries. Orowan considered the boundaries to be regions where dislocations had concentrated. Hess and Barrett&apos; observed the formation of kinks in critically oriented zinc compression specimens. It was found that the sudden snapping over of the tilted region was not an essential part of the kinking phenomenon but rather was associated with the method of loading. Kinking was explained on the basis of the accepted mechanisms of slip and flexural glide. It was also suggested that a kink can be considered as a special type of deformation band. During a recent investigation of the effect of surface condition on creep of zinc single crystals," a similar phenomenon was observed in crystals which were loaded in tension. The specimens were 0.4 in. diam cylindrical rods with a free length of 4 in. between end connections. All of the crystals were tested at a constant load. adjusted to give an extension rate of approximately 0.05 pct per hr. Kinking, as shown in Fig. 1, occurred in some of the tests performed at temperatures above 200°C. Kinking was observed in specimens varying widely in initial orientation, scattering about a 45" angle between the slip plane and the specimen axis. The condition leading to kinking in tension appeared to be non-uniform distribution of flow along the gage length combined with the restraint imposed by the tensile load. At low temperatures and fast strain rates, the distribution of strain throughout the midsection of tension specimens was generally quite uniform; therefore, bend planes developed only near the ends as described by Miller.4 However, at high temperatures under creep conditions it was difficult to obtain specimens in which the distribution of flow was uniform. Some of the factors which may have caused differences in flow stress along the length of the crystals are: Nonuniform distribution of impurities, accidents of growth (lineage structure), slight bending or other damage during handling, surface conditions, and strain-aging characteristics due to dissolved nitrogen." Plastic flow, once started in a local region, often continued to a relatively large strain before other parts of the gage length became active. Under these conditions a series of kinks, such as those in Fig. 1, were produced. Fig. 2 illustrates how a nonuniform distribution of slip produces bending moments which- are responsible for kink formation. If no plastic bending were to occur while the rod extended by pure slip in two separate sections of the rod, it would assume a shape such as that in Fig. 2a. A specimen having this shape and being subjected to a tension load would have concentrations of stress at the concave surfaces C and lower than average stress would exist at the convex surfaces D. Therefore a bending moment is superimposed on the average stress in the regions between C and D. The positions of potential bend planes under these conditions are shown as dotted lines. Actually no such shape as Fig. 2a develops because plastic bending in the region C-D occurs simultaneously with pure slip in the intervening regions. The observed structure of a tension kink is indicated by Fig. 2b. Perhaps the best approach to an understanding of kinking as well as other related plastic bending phenomena, such as the bend planes discussed by Miller,&apos; cell formation studied by Wood et al.,6 and polygonization,7 is consideration of the dislocation model of a bent lattice. Current theories of the slip process postulate generation or multiplication of dislocations at certain lattice imperfections. The mechanism proposed by Frank and Read8 results in continuous generation of concentric dislocation loops which spread out

Jan 1, 1953

By B. C. Raynes

IN the early part of 1952, under the auspices of the U. S. Atomic Energy Commission, Horizons Inc. undertook an investigation dealing with the preparation of high purity thorium metal in order to develop an economically competitive production process for this metal based on a recovery of the metal by a fused salt electrolysis, since a survey of the literature of thorium chemistry indicated that electrolytic methods had not received much attention, despite the consideration that electrochemical production methods often offer attractive commercial potentialities. Thorium metal has been prepared by methods which fall into four general categories: 1) reduction of halides or double halides; 2) reduction of the oxide; 3) thermal decomposition of halides; and 4) electrolytic methods. Some of the important results obtained by various investigators since about 1900 in recorded attempts to prepare pure thorium metal are summarized in Table I. A bibliography of pertinent literature is included in this paper. Thorium is strongly electropositive and possesses a great affinity for oxygen and nitrogen. As a result the prospects for the production of high purity thorium from aqueous media or oxygen bearing compounds are remote. The most promising and profitable area of investigation for production of thorium metal lies in fused salt electrolysis. Thorium has been produced previously by this technique from halide salts in fused alkali chloride eutectic mixtures. Until recently, however, the lack of known large uses for thorium has inhibited the investigation of these methods. The basic problem in an investigation into fused salt production methods for thorium is to develop a suitable fused salt system for the economic preparation of thorium. The first portion of the work in these laboratories was a small scale investigation of the electrolysis of potassium thorium fluoride, KThF5, under a variety of conditions. High purity thorium was prepared in this manner, but the process did not appear to offer the simplest and most attractive continuous commercial operation. The work reported here relates to an all-chloride system employing thorium chloride as the cell feed. ThCL4 was produced as an anhydrous material stabilized in molten sodium chloride. This product, further diluted in NaCl, was subjected to electrolysis. The electrolytic process has been studied in cells capable of holding up to 50 lb total salt charge. Ductile, pure thorium metal has been reproducibly obtained in this operation. Considerable experimentation was carried out in the preparation of ThCL as a cell feed free from moisture, oxides, and other impurities. Thorium was cathodically deposited as a coarsely crystalline powder interspersed with residual bath salts. Aqueous procedures for separating and recovering the metal were developed. The thorium metal powder obtained has been evaluated with respect to its hardness, chemical purity, and certain other physical properties. Preparation of ThC1,—Thorium chloride, and the chlorides of some other metals (aluminum, beryllium, tantalum, zirconium, etc.) cannot be obtained from aqueous solution by direct heating and evaporation. Evaporation and heating of a solution of thorium chloride yields a hydrated salt or partially hydrated salt, and finally Tho,. Anhydrous thorium chloride is obtained by forming a hydrated ammonium chloride complex salt in acidified water solution, dehydrating this compound, and finally decomposing the double salt and subliming the ammonium chloride component. Various methods for the preparation of anhydrous thorium chloride, in accordance with this concept, were investigated as indicated in Table I1 from hydrated thorium chloride (ThCl, . 3H2O), and from thorium oxycarbonate (ThOCO,). In the method found most satisfactory for this work, thorium oxycarbonate was dissolved in an excess of concentrated HC1 to give a clear yellow solution. At least 2 mol NH,Cl were added to the solution per mol of contained thorium. Ammonium thorium chloride hydrate was crystallized from solution by cooling and saturation with HC1. The resultant complex, (NH,Cl),ThCl;xH,O, was dehydrated by evaporation at 130° to 150°C and thoroughly dried at about 250°C. The dried ammanium thorium chloride was then mixed with sodium chloride in proportions to form NaThCL or to produce a ThCL4-NaCl mixture with a given thorium content. The ammonium chloride was sublimed off at temperatures above 500°C in an inert atmosphere furnace to give an anhydrous chloride source electrolyte, essentially free of Tho, and of NKC1.

Jan 1, 1958

By James P. Kimball

Jan 1, 1885

By AIME AIME

THE organization of the Board of Directors was effected Tuesday afternoon in &apos; executive session. Messrs. Bassett,. Buehler, Cates, Daveler, Eavenson, Eilers, Elliott, Evans, Guess, Lovejoy, Moulton, Rickard, Tally, Turner,, Wraith and Wright being pres¬ent. Messrs. Tally, Bassett, Rickard, Mathews and Cates were chosen to form the Executive Committe with Mr. Tally as Chairman, and Messrs. Krumb, Guess and Moulton for the Finance Committee with Mr. Krumb as chairman. Karl Eilers was reelected as Treasurer and H. Foster Bain as Secretary, after which the Directors adjourned for the usual dinner reported elsewhere. At the open session of the Board, at which in addition Messrs. Allen and McAuliffe were present, the work of organization was continued with appointments to committees as listed herewith. Various actions of the Executive and Finance Committees were confirmed, the budget for 1931 formally approved, the meeting dates for the year adopted, 34 new Members and Associates, 13 Junior Members and 31 Student Associates admitted by election and transfer and 29 resignations accepted.

Jan 1, 1931

By E. T. Turkdogan, P. Grieveson

Molten silver is equilibrated with silicon nitride at 1400°C in nitrogen + hydrogen gas mixtures, and from the solubility data the activity coefficient of silicon is found to be 1.76 at silicon concentrations below 1 pct. Using this value, and that available from the silicon cmbide solubility measurements, together with those derived from the phase equilibrium diagvam, the activities of silver and silicon are estimated for the entire composition range. The main feature of the activity data is that there is a pronounced positive deviation from ideality. The activity of silver in molten silver containing up to 5 pct Si was determined by Schadel, Derge, and Birchenall1 by measuring the vapor pressure of pure silver and of the alloy over a temperature range of 850º to 1050°C. They deduced from their results that there is a positive deviation from Raoult&apos;s law and that the activity coefficient of silver increases with decreasing temperature. However, as pointed out by Kubaschewski and Catterall,2 their activity coefficients are much too high, giving silver activities greater than unity. At a later date McCabe and Birchenal13 showed that the previous vapor pressure data on silver4 were high by a factor of about 1.3; presumably a similar error must have occurred with the vapor pressure measurements on the Ag-Si alloys. The activities of silicon and silver calculated by tager&apos; from the phase diagram and by Gibbs-Duhem integration, respectively, show a positive deviation from Raoult&apos;s law for silicon activity and positive and negative deviations for the silver activity. According to these computations, at infinite dilution of silicon ygi = 1.12 at 1420°C. d&apos;Entremont and chipman6 measured the solubility of silicon carbide in molten silver at unit carbon activity; the equilibrium silicon content is found to be Nsi = 0.017 ± 0.002 atom fraction for 1420°C. Using the data of Grieveson and Alcock7 for the free energy of formation of silicon carbide from pure silicon and graphite, the activity of silicon for unit activity of carbon and of silicon carbide is asi = 0.0308 at 1420°c; therefore for the above solubility data on silver, YSi = 1.81, which is somewhat higher than that given by Hager. Use is often made of silver alloys in the measurements of activities of solutes in metals where silver has a very low solubility; in fact, Ag-Si and Ag-A1 alloys have been used in the past for such purposes. In an attempt to gain more thermodynamic data on molten Ag-Si alloys, measurements have been made in the present study on the solubility of silicon nitride in molten silver; the results on these solubilities are given in this paper together with a detailed discussion of activities of silver and silicon. SOLUBILITY OF SILICON NITRIDE The standard free energy of formation of silicon nitride is known with a reasonably high degree of accuracy. According to the measurements made by Hincke and Brantle8 and by Pehlke and Elliott,&apos; the following are the free energy data for the standard states of solid Si3N4, liquid silicon, and Nz at 1 atm pressure:

Jan 1, 1963

By Howard N. Eavenson

Waste of coal, or perhaps more properly the percentage of its recovery in mining, has keenly interested me during an experience of over a half century in coal mining. In the early part of that time and until a few years after World War I the use of better mining plans and supervision resulted in an increased recovery. Since then, and particularly since controlled prices have been effective and mechanized mining has been extensively practiced, the tendency has all been toward lesser recovery in mining. The purpose of this paper is not to criticize the methods being used, which after all are the result of economic forces, nor to point to any particular mining method or to any locality where the best recovery is obtained, but to call attention to the tremendous waste of our national coal resources that is constantly continuing. A possible method of reducing this loss will be suggested, but without any illusions that it will be followed or that such attention will be paid to it. Average Recovery The actual average percentage of recovery of coal in the ground is not known. Many papers giving figures of recovery have been published but no definite base of measurement has been agreed upon by those presenting the data. As an example, the Pittsburgh bed can be considered—it has produced a greater output than any other and has been sold or leased at higher prices in place than any other coal bed. Over a large part of its area where high prices have prevailed it has been customary for many years to draw pillars. In these mines it has been usually assumed that the average recovery was higher than in other beds and percentages of recovery of 85 to 95 per cent are usually said to be realized. In all of these mines, however, it has been customary to leave from 4 to 8 in. in the bottom, and 10 to 16 in. in the roof; thus from 14 to 24 in. of coal has been left in the mines out of a total seam thickness of about 96 in., or 15 to 25 per cent of the total coal. No attempt is made to mine this coal for metallurgical purposes, owing to its impurities, although this unmined coal is of excellent quality and better than much of the coal used for steam purposes. If we add to this discard the coal usually lost in mining and those parts of the bed &apos;that must be left because of bad seam conditions, barrier pillars, reservations for buildings, gas and oil wells, one can easily see that the actual recovery from this bed cannot be more than 65 to 75 per cent. In addition, the rooster or rider or wild coal over the mining bed—often 3 to 4 ft. of coal, high in ash and sulphur but still good steam coal—is hardly ever mined, nor is any attempt made to mine it. It is safe to say that from a broad economic viewpoint not more than 50 to 60 per cent of the fuel value of this great seam is actually being recovered.

Jan 1, 1947

By Howard N. Eavenson

Waste of coal, or perhaps more properly the percentage of its recovery in mining, has keenly interested me during an experience of over a half century in coal mining. In the early part of that time and until a few years after World War I the use of better mining plans and supervision resulted in an increased recovery. Since then, and particularly since controlled prices have been effective and mechanized mining has been extensively practiced, the tendency has all been toward lesser recovery in mining. The purpose of this paper is not to criticize the methods being used, which after all are the result of economic forces, nor to point to any particular mining method or to any locality where the best recovery is obtained, but to call attention to the tremendous waste of our national coal resources that is constantly continuing. A possible method of reducing this loss will be suggested, but without any illusions that it will be followed or that such attention will be paid to it. Average Recovery The actual average percentage of recovery of coal in the ground is not known. Many papers giving figures of recovery have been published but no definite base of measurement has been agreed upon by those presenting the data. As an example, the Pittsburgh bed can be considered—it has produced a greater output than any other and has been sold or leased at higher prices in place than any other coal bed. Over a large part of its area where high prices have prevailed it has been customary for many years to draw pillars. In these mines it has been usually assumed that the average recovery was higher than in other beds and percentages of recovery of 85 to 95 per cent are usually said to be realized. In all of these mines, however, it has been customary to leave from 4 to 8 in. in the bottom, and 10 to 16 in. in the roof; thus from 14 to 24 in. of coal has been left in the mines out of a total seam thickness of about 96 in., or 15 to 25 per cent of the total coal. No attempt is made to mine this coal for metallurgical purposes, owing to its impurities, although this unmined coal is of excellent quality and better than much of the coal used for steam purposes. If we add to this discard the coal usually lost in mining and those parts of the bed &apos;that must be left because of bad seam conditions, barrier pillars, reservations for buildings, gas and oil wells, one can easily see that the actual recovery from this bed cannot be more than 65 to 75 per cent. In addition, the rooster or rider or wild coal over the mining bed—often 3 to 4 ft. of coal, high in ash and sulphur but still good steam coal—is hardly ever mined, nor is any attempt made to mine it. It is safe to say that from a broad economic viewpoint not more than 50 to 60 per cent of the fuel value of this great seam is actually being recovered.

Jan 1, 1947

By Howard N. Evanenson

WASTE of coal, or perhaps more properly the percentage of its recovery in mining, has keenly interested me during an experience of over a half century in coal mining. In the early part of that time and until a few years after World War I the use of better mining plans and supervision resulted in an increased recovery. Since then, and particularly since controlled prices have been effective and mechanized mining has been extensively practiced, the tendency has all been toward lesser recovery in mining. The purpose of this paper is not to criticize the methods being used, which after all are the result of economic forces, nor to point to any particular mining method or to any locality where the best recovery is obtained, but to call attention to the tremendous waste of our national coal resources that is constantly continuing. A possible method of reducing this loss will be suggested, but without any illusions that it will be followed or that such attention will be paid to it. AVERAGE RECOVERY The actual average percentage of recovery of coal in the ground is not known. Many papers giving figures of recovery have been published but no definite base of measurement has been agreed upon by those presenting the data. As an example, the Pittsburgh bed can be considered-it has produced a greater output than any other and has been sold or leased at higher prices in place than any other coal bed. Over a large part of its area where high prices have prevailed it has been customary for many years to draw pillars. In these mines it has been usually assumed that the average recovery was higher than in other beds and percentages of recovery of 85 to 95 per cent are usually said to be realized. In all of these mines, however, it has been customary to leave from 4 to 8 in. in the bottom, and 10 to 16 in. in the roof; thus from 14 to 24 in. of coal has been left in the mines out of a total seam thickness of about 96 in., or 15 to 25 per cent of the total coal. No attempt is made to mine this coal for metallurgical purposes, owing to its impurities, although this unmined coal is of excellent quality and better than much of the coal used for steam purposes. If we add to this discard the coal usually lost in mining and those parts of the bed that must be left because of bad seam conditions, barrier pillars, reservations for buildings, gas and oil wells, one can easily see that the actual recovery from this bed cannot be more than 65 to 75 per cent. In addition, the rooster or rider or wild coal over the mining bed often 3 to 4 ft. of coal, high in ash and sulphur but still good steam coal-is hardly ever mined, nor is any attempt made to mine it. It is safe to say that from a broad economic viewpoint not more than 50 to 60 per cent of the fuel value of this great seam is actually being recovered.

Jan 1, 1946

By Forrest K. Pence

BENTONITE deposits are known to occur in Texas within the Jackson group of formations. This group represents the uppermost Eocene age sediments found in the coastal plain area of Texas. It outcrops across this area of the state in a narrow band of some 4 to 20 miles width. The outcrop pattern roughly parallels the present Gulf of Mexico shore line and is some 100 miles inland from the Texas shore, Fig 1. The principal bentonite deposits are found in the areas where this outcrop pattern cuts across the south-central Texas counties of Karnes, Gonzales, and Fayette. In these deposits, the better quality bentonite is found in the lower or bottom layers of the volcanic ash deposits in which they occur. Some of these better quality bentonites develop very light colors upon firing and therefore justify their being classified as "white firing." The deposits in Karnes and Gonzales Counties apparently occur in commercial quantity, whereas the white firing strata so far uncovered in Fayette County have been too thin to be classified as yet as "commercial." A study of the ceramic properties of the weathered ash in Gonzales and Karnes Counties was reported in 1941.1

Jan 1, 1949

By Stanley M. Moos, Richard Addison

Introduction The Pueblo Viejo gold-silver mine, located in the Dominican Republic, started production in early 1974 processing ores averaging 4 g/t gold and 20 g/t silver at a rate of 7.25 kt/d. The ore milled is a silicified siltstone with average work index of about 8 that occurs as an extensive oxide capping of a primary sulfide ore body. The sulfide ores are refractory to conventional treatment process and extensive research work is now in progress to develop a treatment process for them. The ore deposit has a long history, having been mined by the indigenous Arawak Indians for some time in the early sixteenth century prior to the arrival of the Spanish, and then by the Spaniards. In the early 1950s, Italian engineers attempted to process the sulfide ores. Mining by the Indians and Spanish was abandoned because of low grade and little coarse gold in the oxide ores and by the Italians because of the refractory nature of the sulfide ore. In 1969, Rosario Resources geologists discovered that there existed sufficient reserves of oxide ore for a viable mining operation and the plant was designed, built, and brought into production.

Jan 1, 1981