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By R. J. Brison, O. F. Tangel

The development of a new method of mineral separation was sponsored by the International Salt Company, which requested Battelle Institute to investigate means for improving the quality and appearance of rock salt from the Company's Detroit mine. Although developed specifically for removing impurities from rock salt, the general method may be applicable to other separation problems. The principal impurities in rock salt from the Detroit mine are dolomite and anhydrite which represent 2 to 5 pct of the weight of the mined salt. In the size range from 1/4 to M in. (the range of primary interest in this project) the impurities are only partially liberated from the halite in normal production. Further size reduction to improve the liberation of impurities is not practicable in view of the market requirements for the coarse grades of rock salt. Laboratory separations in heavy liquids showed that, to improve the quality and appearance of the rock salt substantially, it would be necessary to remove not only free gangue particles but also a large proportion of the locked-in particles. Because rock salt is an inexpensive commodity, a low-cost process was required. Gravity methods were, of course, considered. The heavy-liquid separations indicated that a split at an effective specific gravity of 2.2 to 2.3 would be required. (The specific gravity of pure halite is 2.16.) Heavy-media separation was investigated but had the disadvantages that it was necessary both to operate with saturated brine and to dry the cleaned salt, and that the cleaned salt was darkened by the magnetite medium. Air tabling was tried but did not give the desired separation. It soon became apparent that established methods would not provide a satisfactory solution and work was undertaken on the development of a new process to solve the problem. PROCESS DEVELOPMENT Preliminary Experiments: At the start of the investigation, an analysis of the problem indicated that the diathermacy of rock salt—that is, its ability to transmit radiant heat—might form the basis for an efficient separation process. Under this theory, the impurities might be selectively heated by radiant heat. The particles could then be fed over a belt coated with a heat-sensitive substance so that the warm impure particles would adhere preferentially to the coating. After the initial experiments, made by heating the rock salt with an infrared lamp and separating the product on small sheets of resin-coated rubber, proved encouraging, a small continuous separation unit was set up. This comprised 1) a simple heating unit consisting of a vibrating feeder covered with aluminum foil and an infrared lamp mounted above the feeder and 2) a separation belt 6 in. wide and 36 in. long. A sketch of the device is shown in Fig. 1. Results with this apparatus confirmed the fact that a good separation was possible. It was apparent, however, that a considerable amount of experimental work would be needed to develop the scheme to a practical and economical process. The Process: Basically, the process consists of two main steps: 1) selective heating by radiation and 2) separation of the heated particles on a heat-sensitive surface. Because neither of these steps had previously been utilized commercially in mineral processing, it was necessary to do basic research on both aspects. Factors studied in the investigation included type of heat source, design of heating unit, design of separation belt, selection of heat-sensitive coating, removal of heated particles from the belt, contact between particles and coating, and maintenance of the heat-sensitive surface. Part of the experimental work was carried out on a small-scale unit consisting of the 36x6 in. belt and auxiliary apparatus, and part on a larger unit. For simplicity, discussion of work on both of these units is grouped together. SELECTIVE HEATING Radiant-Heat Source: The essential requirements for a radiant-heat source were 1) that the radiant heat be in a wave length range which is effectively absorbed by the impurities but not absorbed appreciably by the rock salt and 2) that it be dependable, practical, and economical. Selection of a heat source of suitable wave length range was one of the first considerations. It is well known that pure halite is highly transparent to radiant energy in wave lengths from 0.3 to 13 microns. However, the available data on infrared transmission by dolomite and anhydrite, particularly in the range below two microns, were not complete enough to serve as a reliable basis for selection of a heat source. Although it may have been possible to obtain sufficient data on infrared transmission and absorption to enable one to select the best heat source, a more direct procedure was used. This consisted simply of exposing the crude rock salt to each of several types of radiant-heat source on the small continuous separation device. The heat sources investigated, approximate source temperature used, and calculated wave length of maximum radiation are tabulated in Table I. Of the two types of tungsten-filament lamps investigated, both the short wave length photoflood lamps and the longer wave length infrared lamps were satisfactory from the standpoint of selectivity

Jan 1, 1961

By R. L. Dietrich

Magnesium alloy sheet has less ability to accept bending at room temperature than most of the heavier metals. In work designed to improve the bend properties, the preferred orientation of the sheet is of major importance as it is in all studies of the properties of wrought magnesium products. When rolled into sheet, all of the common magnesium alloys form an orientation texture having the basal (002) planes approaching parallel to the surface of the sheet. This texture is only slightly affected by annealing. Magnesium single crystals are highly anisotropic, and, as might be expected, so are magnesium alloy wrought products in which a strong preferred orientation is developed. It is therefore not surprising that bend properties are affected by orientation. Ansel and Betterton1 reported that the orientation of AZ3lXt sheet varies from surface to center and that bend properties are improved by etching away the sharply oriented material at the surface of the sheet to reach the more broadly oriented structure below. This paper covers a study of that orientation, either during the rolling process or by treatment of the finished sheet, in an effort to improve the bend properties and toughness of sheet. Literature The orientation texture of magnesium and magnesium alloy sheet has been studied extensively. Early determinations2 showed that pure magnesium sheet has a preferred orientation in which the basal planes are parallel to the sheet surface within very narrow limits. J. C. hIcDonald3 and J. D. Hanawalt4 reported that sheet containing a small amount of calcium develops a "double" texture, that is, the majority of the basal planes are a few degrees from parallel to the surface and there is a noticeable vacancy at the parallel position. Bakarian5 made careful quantitative pole figures of both pure magnesium sheet and MI alloy SEPTEMBER 1949 sheet which show these features. In all of these studies, however, the orientation was determined by transmission methods in which the resulting pattern is an average through the thickness of the sheet. The tendency of wrought metal to exhibit a different orientation at the surface from that in the center has been reported by many investigators. G. von Vargha and G. Wasserman6 found that with copper, aluminum, iron, and brass the textures of rolled compared to drawn wires were the same at the center but differed markedly at the surface. It was also reported by investigators7 that the orientation of rolled aluminum varies from surface to center. Har-greaves8 found that the surface texture of AM503 (magnesium alloy similar to MI) sheet was different from the center texture. It is reported by Edmunds and Fuller9 that zinc alloy sheet sometimes had a thin layer at the surface with a strong orientation of the basal planes parallel to the surface, which, if present, impaired the bend properties of the sheet. Part1 Surface Orientation ofMag- nesium Alloy Sheet and the Effect on Properties Attempts to correlate the bend properties of magnesium alloy sheet with tension ductility over short gauge lengths proved unsuccessful and the subsequent investigation showed that nonuniformity in orientation is a con- tributing factor as the properties of the surface material have a much more important effect in bending than in tension. A program to study the relationship between surface orientation at the surface and bend properties was then undertaken. First, the effect of etching away the surface of sheet on the bend properties and the orientations at the various depths were studied. Sheet samples of M1, AZ31X, and AZ61X were etched in dilute nitric acid to remove the surface material for various depths to 0.015 in. As may be seen in Table 1, the minimum bend radius improved considerably as the surface layers were etched away but it was necessary to etch the sheet quite deeply, much more so than was found necessary by Edmunds and Fuller9 on zinc sheet. It is also apparent that the amount of etching required is a function of the sheet thickness. In all of this work, radii were measured as R/t, the radius divided by the sheet thickness, in order to eliminate the effect of the reduction in sheet thickness produced by the etching. To determine the orientation texture of the sheet, X ray reflection patterns were taken using copper radiation with the bearn striking the specimen at an angle of 17' to the surface, which is the Bragg angle for the (002) planes of magnesium. Two exposures were made of each specimen, one with the beam perpendicular to the rolling direction and the other with the beam parallel to the rolling direction. The symmetry of the preferred orientation in magnesium sheet is such that these two photographs gave an approximation of the pole figure sufficiently accurate for qualitative work and it was not thought worthwhile to make complete pole figures. These X ray patterns show that the orientation has a much narrower spread at the original surface of the sheet than below the surface. The narrow spread is found in sheet having the majority of the basal planes (002) parallel to the surface, and since this is an unfavorable position for slip, it is

Jan 1, 1950

By A. Y. Bethune, W. G. Woolf

THE hydrometallurgy of special high-grade zinc as practiced by the Sullivan Mining Co. at its electrolytic zinc plant, Kellogg, Idaho, involves an important filtration step immediately following the leaching process. By means of the filtration the heavy zinc sulphate solution is separated from the residual products which remain after the zinc calcine has been dissolved in the sulphuric acid electrolyte. Because this plant uses the so-called high-acid, high-density process' for the production of First, the strength of the electrolyte (270g H,SO, per liter) results in a saturated zinc sulphate solution, having a specific gravity of 1.510 to 1.540, which must be kept warm during filtration because of its property of "seeding out" small crystals if allowed to drop much below 60°C. Second, the action of the "high" acid on zinc calcine under the temperature conditions of the leach (80" to 102 "C), although favorable to good zinc extraction, causes a considerable quantity of iron to be dissolved (8 to 18. g per liter) along with variable quantities of alumina and silica, depending on the grade and type of original zinc concentrates roasted. These three, iron, alumina, and silica, are almost completely precipitated during the neutralization of the leach (only a few. milligrams per liter of each remain in solution), so that the resulting pulp, instead of being a granular, sand-like product having a particle-size distribution dependent on the fineness of the zinc calcines leached, is in reality a slimy, chemical precipitate whose filtration characteristics constantly change depending on the amounts of iron silica, and other impurities, which are dissolved and reprecipi-tated. Third, the combination of supersaturated solution of high specific gravity plus a dense, semi-gelatinous residue creates a difficult washing problem requiring a positive displacement wash to liberate the zinc sulphate entrapped in the pulp. In a closed-cycle hydrometallurgical operation, such as practiced in this plant, the extent of washing is determined by the volum,e limitations imposed on the intermediate wash waters by the amount of "fresh" (or process) water which may be added. The volume of fresh water used for makeup purposes is limited to the amount which is lost during the closed cycle by evaporation in the leach, sulphate content of the calcines leached, moisture content of the residue, and spillage. The Burt filter as modified and improved by the Sullivan Mining Co. has successfully met and overcome these difficulties under a variety of zinc plant operating conditions since 1928. It might have many interesting applications to metallurgical fields other than that of electrolytic zinc, and its possible usefulness to hydrometallurgists in general warrants its description and discussion. The Burt filter is so named from its inventor who originated it in Mexico for pulp filtration in the cyanide process for gold and silver ores. While retaining the basic principle of Burt's earlier revolving pressure-type filter with internal filtration media, a number of modifications and improvements have been made in Sullivan Mining Co.'s installation. The Burt filter may be classified as a batch-type pressure filter in contradistinction to either the conventional vacuum-type filter, which depends on atmospheric pressure to force solution through a cloth medium, or to the filter-press, which employs whatever pressure is imparted by the pump delivering the liquid being filtered. The Burt consists essentially of a hollow steel cylinder about 40 ft long, 5 ft in diameter, resting horizontally, and capable of rotation about its long axis. It is supported on one end by a hollow trunnion and near the other end by a riding-ring and roller combination. The cylinder is lined with filter units each fastened against the inside of the shell and parallel to the long axis so as to form a hollow cavity into which pulp may be charged. A specific amount of pulp is admitted to the filter and a unique valving arrangement prevents the loss of pulp while air pressure forces the solution through a canvas medium to the discharge port of each filter unit. The residue is left on the surface of the canvas inside the cavity. The remainder of the filter cycle is concerned with washing the residue free of zinc sulphate, discharging it from the Burt, and preparing the filter for the next charge. A more detailed description of Burt filter construction, a typical filter cycle, and its operating characteristics when employed on material encountered in this plant will be given in that order. Description of the Filter: Fig. 1 shows a side elevation view of a filter with riveted shell construction. Since this drawing was made shells have been fabricated by welding, instead of riveting, with complete success. Shells are lagged on the outside to retain heat. Fig. 1 shows a side elevation and plan view of a Burt filter in operating position. The 1/2-in. steel shells are lined with 3/16-in. copper sheet as protection against the corrosive action of the solution (containing about 500 mg Cu per liter) on iron, and the copper is given a thin protective coating of plastic-base paint. Fig. 2 is a view from the discharge end of the filter, with head removed, before filter units are fastened to the periphery. It shows

Jan 1, 1951

By W. G. Woolf, A. Y. Bethune

THE hydrometallurgy of special high-grade zinc as practiced by the Sullivan Mining Co. at its electrolytic zinc plant, Kellogg, Idaho, involves an important filtration step immediately following the leaching process. By means of the filtration the heavy zinc sulphate solution is separated from the residual products which remain after the zinc calcine has been dissolved in the sulphuric acid electrolyte. Because this plant uses the so-called high-acid, high-density process' for the production of First, the strength of the electrolyte (270g H,SO, per liter) results in a saturated zinc sulphate solution, having a specific gravity of 1.510 to 1.540, which must be kept warm during filtration because of its property of "seeding out" small crystals if allowed to drop much below 60°C. Second, the action of the "high" acid on zinc calcine under the temperature conditions of the leach (80" to 102 "C), although favorable to good zinc extraction, causes a considerable quantity of iron to be dissolved (8 to 18. g per liter) along with variable quantities of alumina and silica, depending on the grade and type of original zinc concentrates roasted. These three, iron, alumina, and silica, are almost completely precipitated during the neutralization of the leach (only a few. milligrams per liter of each remain in solution), so that the resulting pulp, instead of being a granular, sand-like product having a particle-size distribution dependent on the fineness of the zinc calcines leached, is in reality a slimy, chemical precipitate whose filtration characteristics constantly change depending on the amounts of iron silica, and other impurities, which are dissolved and reprecipi-tated. Third, the combination of supersaturated solution of high specific gravity plus a dense, semi-gelatinous residue creates a difficult washing problem requiring a positive displacement wash to liberate the zinc sulphate entrapped in the pulp. In a closed-cycle hydrometallurgical operation, such as practiced in this plant, the extent of washing is determined by the volum,e limitations imposed on the intermediate wash waters by the amount of "fresh" (or process) water which may be added. The volume of fresh water used for makeup purposes is limited to the amount which is lost during the closed cycle by evaporation in the leach, sulphate content of the calcines leached, moisture content of the residue, and spillage. The Burt filter as modified and improved by the Sullivan Mining Co. has successfully met and overcome these difficulties under a variety of zinc plant operating conditions since 1928. It might have many interesting applications to metallurgical fields other than that of electrolytic zinc, and its possible usefulness to hydrometallurgists in general warrants its description and discussion. The Burt filter is so named from its inventor who originated it in Mexico for pulp filtration in the cyanide process for gold and silver ores. While retaining the basic principle of Burt's earlier revolving pressure-type filter with internal filtration media, a number of modifications and improvements have been made in Sullivan Mining Co.'s installation. The Burt filter may be classified as a batch-type pressure filter in contradistinction to either the conventional vacuum-type filter, which depends on atmospheric pressure to force solution through a cloth medium, or to the filter-press, which employs whatever pressure is imparted by the pump delivering the liquid being filtered. The Burt consists essentially of a hollow steel cylinder about 40 ft long, 5 ft in diameter, resting horizontally, and capable of rotation about its long axis. It is supported on one end by a hollow trunnion and near the other end by a riding-ring and roller combination. The cylinder is lined with filter units each fastened against the inside of the shell and parallel to the long axis so as to form a hollow cavity into which pulp may be charged. A specific amount of pulp is admitted to the filter and a unique valving arrangement prevents the loss of pulp while air pressure forces the solution through a canvas medium to the discharge port of each filter unit. The residue is left on the surface of the canvas inside the cavity. The remainder of the filter cycle is concerned with washing the residue free of zinc sulphate, discharging it from the Burt, and preparing the filter for the next charge. A more detailed description of Burt filter construction, a typical filter cycle, and its operating characteristics when employed on material encountered in this plant will be given in that order. Description of the Filter: Fig. 1 shows a side elevation view of a filter with riveted shell construction. Since this drawing was made shells have been fabricated by welding, instead of riveting, with complete success. Shells are lagged on the outside to retain heat. Fig. 1 shows a side elevation and plan view of a Burt filter in operating position. The 1/2-in. steel shells are lined with 3/16-in. copper sheet as protection against the corrosive action of the solution (containing about 500 mg Cu per liter) on iron, and the copper is given a thin protective coating of plastic-base paint. Fig. 2 is a view from the discharge end of the filter, with head removed, before filter units are fastened to the periphery. It shows

Jan 1, 1951

By William M. Boorstein, Robert D. Pehlke

The rates of solution (of hydrogen in liquid pure iron and in several liquid binary iron alloys were meas-ured using a constant volume technique. The rates of absorption and desorption were found to be equal un-der all experimental conditions. increasing concen-trations of S, Si, or Te decrease the rate of hydrogen uptake but additions of Al, B, Cr, Cu, or Ni have no measurable effect up to concentrations normally en-countered in steelmaking practice. No relation ship was found between the effect of an alloying element on the equilibrium solubility of hydrogen in liquid iron and its effect on the solution rate constant. Mathe-rnatical analysis of the data indicates that under the present experimental conditions the rate of reaction of hydrogen with liquid iron is controlled by transport of gas solute atoms in the metal phase. Comparison of the present resuts with data on nitrogen taken un der similar conditions establishes that the hydrody-nurnic conditions which exist near the surface of a metal bath are best approximated mathematically by a surface renewal model for the case of rapid in-ductive stirring and by a boundary layer model for more quiescent melts. HYDROGEN has long been recognized as being a detrimental constituent in steel. If dissolved in the molten metal in excess of its solid solubility, hydro-gen can be evolved during solidification and cause bleeding or porosity in ingots and castings. In the solid metal, lesser amounts play a definite role in causing other defects such as hairline cracks, blisters, and embrittlement. For significant refinements to be made in metallurgical procedures designed to control or eliminate hydrogen from liquid iron or steel dur-ing processing, available equilibrium solubility data must be supplemented with reliable fundamental in-formation pertaining to the kinetic factors involved in the transfer of hydrogen to or from the metal. The scarcity of such information in the literature prompted the present investigation. PREVIOUS RESEARCH Whereas much of the existing data on the solution kinetics of gases such as nitrogen were obtained during the course of thermodynamic investigations, the solu-tion rate of hydrogen has been found too rapid to be accurately determined by conventional solubility meas-urement techniques. Consequently, little work on hy-drogen solution kinetics has been reported in the lit-erature. Carney, Chipman, and crant1 attempted to study the rate of solution and evolution of hydrogen from liquid iron by employing a newly devised sampling method. Although no significant quantitative data could be obtained, it was observed that the rate of solution was approximately equal to the rate of evolution of hy-drogen from the melt. Karnaukov and Morozov2 stud-ied the rate of absorption and Knuppel and Oeters3 the rate of desorption of hydrogen from molten iron by measuring pressure changes with time in a constant volume system. Karnaukov and Morozov determined the hydrogen pressures over their inductively stirred melts with the aid of a McLeod gage and therefore, were forced to work at pressures not in excess of 40 mm of Hg. Their experimental data conformed to a mathematical correlation based on diffusion control: and the rate coefficients calculated on this basis were shown to be independent of the initial absorption pres-sure. These authors reported the solution rate of hy-drogen to be eight-to-ten times higher than they had found for nitrogen in a previous study. They also re-ported that under identical conditions, hydrogen dis-solves somewhat more slowly in iron-columbium alloys than in pure iron. Knuppel and Oeters found that the desorption of hydrogen from pure iron at 1600°C was controlled in all cases investigated by diffusion in the metal bath as long as bubble formation was sup-pressed. This was substantiated by Levin, Kurochkin, and umrikhin4 who studied the kinetics of hydrogen evolution from liquid (technical) iron while applying a vacuum. Salter5 measured the rate of hydrogen ab-sorbed by iron buttons, arc-melted by direct current, as a function of hydrogen partial pressure in a hy-drogen-argon atmosphere. A carrier gas technique was used for analysis of the hydrogen absorbed. The initial rate of absorption was found to increase di-rectly with the square root of the partial pressure of hydrogen. EXPERIMENTAL METHOD Because of the rapid uptake and evolution of hydro-gen by iron-base melts, a constant volume technique was devised in order to obtain meaningful kinetic data over the entire course of the solution process. Apparatus. A schematic view of the experimental apparatus is given in Fig. 1. The hydrogen-liquid iron reaction system consisted of a gas storage bulb con-nected to a meltcontaining reaction chamber through a normally-closed solenoid valve. The gas storage bulb, an inverted 250 ml round-bottomed Pyrex flask was joined to the inlet port of the solenoid valve by a glass-to-metal seal. A more detailed illustration of the reaction chamber is shown in Fig. 2. The design of the Vycor reaction bulb was essentially that de-scribed by Weinstein and Elliott6 with the exception of a shorter, larger diameter gas inlet for this kinetic study. In position, the reaction bulb was closely by an eight-turn coil of water-cooled copper tubing which, when energized by a 400-kc oscillator, provided the inductive heating source. The walls of the bulb were maintained relatively cool by circulating cold water along their outer surface, thus preventing

Jan 1, 1970

By R. F. Dickerson, D. A. Vaughan, A. F. Gerds

ALTHOUGH the metallurgy of uranium has been under intensive study since the early 1940's, no systematic effort has been made to identify the non-metallic inclusions in uranium. Uranium carbide (UC), which is probably the most common inclusion found in graphite-melted metal, has been tentatively identified by previous investigators, but the other nonmetallic inclusions have received little attention. Since metallography is a valuable tool in metallurgical studies, the metallographic identification of the nonmetallic inclusions in uranium is important. Such an investigation has been completed and the identification of slag-type inclusions and of uranium monocarbide, uranium hydride, uranium dioxide, uranium monoxide, and uranium mononitride is described. Metallographic Preporation It is often possible to prepare specimens for metal-lographic examination equally well by several methods. The specimens which were examined in this work were prepared by one of two acceptable methods. For the convenience of the reader, both methods will be discussed in detail and will be referred to simply as Method I or Method II in the subsequent sections. For both Methods I and 11, specimens for microscopic examination usually were mounted either in bakelite or in Paraplex room temperature mounting plastic. Method I—Specimens were ground in a spray of water on a revolving disk covered successively with 120-, 240-, and 600-grit silicon carbide papers. It was necessary to perform the final grinding operation carefully on worn 600-grit paper to keep the scratches as fine as possible. After washing and drying, the specimens were polished for 3 to 4 min on a slow speed wheel (250 rpm) covered with a medium nap cloth. Diamet Hyprez Blue diamond polishing paste, Grade 00, 0 to 2 µ, was used as abrasive with kerosene as lubricant on the wheel. Specimens were washed thoroughly in alcohol and final polished electrolytically in an electrolyte composed of 1 part stock solution (118 g CrO, dissolved in 100 cm3 H2O) with 4 parts of glacial acetic acid. A stainless steel cathode was used. At an open circuit potential of 40 v dc, a polishing time of 2 sec retained inclusions well with the bath at room temperature. If additional etching was required to sharpen the interface between the metal and the inclusions, an electrolyte composed of 1 part stock solution (100 g CrO3 and 100 cm8 H20) and 18 parts glacial acetic acid was used at room temperature. Best results were obtained by etching for from 10 to 15 sec at 20 v dc in the open circuit. Surfaces obtained by this method are suitable for microscopic examination. However, if desired, they may be etched further with other chemicals. Method 11—Rough grinding was done on a wet 180- or 240-grit continuous grinding belt. The specimen was then ground by hand successively on 240-, 400-, and 600-grit silicon carbide papers in a stream of water. Final polishing was accomplished on a 4 in. high speed wheel (3400 rpm) covered with Forstmann's cloth. Linde B levigated alumina, suspended in a 1 volume pet chromic acid solution, was the abrasive. Specimens usually were polished in 5 min or less by this technique. Often the inclusions present in the metal were identified in the mechanically polished condition. When etching was required to outline inclusions more sharply, one of the two following methods was used. In the first method, the specimen is etched lightly while electropolishing in the chromic-acetic acid solution described above (1 part of stock solution to 4 parts of acetic acid). The electrolyte was refrigerated in a dry ice-ethyl alcohol bath and specimens were etched at 60 v dc on the open circuit for 2 or 3 cycles of 3 to 4 sec each. The second technique utilizes electrolytical etching at about 10 v dc (open circuit) in a 10 pet citric acid solution at room temperature. X-Ray Diffraction Technique The major problem in the identification of inclusions in metals by X-ray diffraction techniques is the extraction of a sufficient amount of each type of inclusion to obtain an X-ray diffraction pattern. In the present study, X-ray diffraction patterns were obtained from individual inclusions of the order of 10 µ diam. The polished and etched samples shown in the micrographs were examined at a magnification of X54 or XI00 with a binocular microscope. This allowed sufficient working distance to extract the inclusions with a needle probe for powder X-ray diffraction analysis. Friable inclusions such as MgF2, CaF2, UO2, and UH3 could be freed from the metal by probing the as-polished and etched surface. The fine particles then were picked up on the end of a Vistanex-coated glass rod (0.002 in. diam) which was held in a brass adapter made to fit the powder X-ray diffraction camera. The end of the glass rod was centered in the path of the X-ray beam. In the case of the UC, UO, and UN inclusions which are smaller in size, more metallic in appearance, and less friable than the other inclusions, it was necessary to etch the inclusion in relief before extraction. UN inclusions etched sufficiently in relief in the electrolytic polishing solution described in Methods I and II by increasing the polishing time. UN inclusions were relief etched by extending the

Jan 1, 1957

By S. Sumski, M. B. Panish, R. T. Lynch

The liquidus isotherms in the gallium-rich corner of the Ga-Al-P phase diagram have been determined from 1000" to 1200°C and at I100°C the corresponding solidus isotherm was obtained. A simple thermody-namic treatment which permits calculation of the solidus and liquidus isotherms is discussed. A technique which was previously used for the growth of GaxAl1-xAs was used for the preparation of solution epitaxial layers of GaxAl1-xP and ALP. An approximate value of 2.49 i 0.05 ev for the band gap of Alp at 300°K was obtained and the ternary phase data were used to estimate a value of 36 kcal per mole for the heat of formation 0f Alp at that temperature. The Gap-A1P crystalline solid solution is one in which there exists the possibility of obtaining crystals with selected energy gaps, within the limits imposed by the energy gaps of Gap and Alp. Such crystals are of considerable interest because of their potential value for optoelectronic and other solid-state devices. Furthermore, it has been amply demonstrated for GaAs and GaP,'-7 that device, or bulk materials grown from gallium solution generally have more efficient radiative recombination than materials prepared in other ways. This presumably due to the lower gallium vacancy concentration in such material.= Small crystals of GaXAl1-xP and A1P have been grown from solution,8-10 and A1P has been grown from the vapor," but neither have previously been grown by liquid epitaxy. In this paper we present the ternary liquidus-solidus phase diagram of the Ga-A1-P system in the region of primary interest for solution epitaxy, and discuss the thermodynamic implications of that phase diagram with particular reference to the liquidus and solidus isotherms in the gallium-rich corner of the GaxAl1-xP primary phase field and to the A1-P system. Several measurements of the absorption edge of GaxAl1-xP crystals have been made and the width of the forbidden gap of A1P has been estimated from these measurements. EXPERIMENTAL The differential thermal analysis technique used to determine the liquidus isotherms and the optical measurements used in this work are similar to those described previously12 for the Ga-Al-As system, ex- thermocouples in the thermopile for added sensitivity. The materials used were semiconductor grade Ga, Gap, and Al+ The composition and temperature range at which DTA studies could be done was quite restricted. The upper temperature was limited by the chrome l-alumel thermopile to about 1200°C, and the highest aluminum concentration to about 5 at. pct by low sensitivity caused by the reduced solubility of Gap with increasing aluminum concentration in the liquid. DTA studies were not possible at 1000°C and below because of the low sensitivity caused by low solubility of Gap in the Ga-A1-P system. The cooling rate for these studies was about 1°C per min. No heating studies were done because of limited sensitivity. Supercooling probably does occur, but our experience with other 111-V systems indicates that it is no greater than about 10 to 15.c. Solid solubilities were determined by analyzing epitaxial layers of GaxAl1-xP grown from the liquid, with an electron beam microprobe. The layers were grown on Gap seeds by a tipping technique in which the layer is grown over a short-temperature range (20" to 50°C) on the seed from a solution of known composition. The tipping technique reported by Nelsson1 for GaAs could not be used, particularly for solutions containing appreciable amounts of aluminum, because of the formation of an A1203 scum on the liquid surface. A system was therefore designed, which would effectively remove the oxides mechanically, so that uniform wetting and crystal growth could occur. This tipping technique has already been described in detail." The best control over the composition of the re-grown layer was obtained when the tipping was done at a temperature which corresponded to the temperature of first formation of solid for the solution being used. Generally, therefore, a solution was prepared by adding the amounts of Ga, Gap, and A1 required to yield a solution which would be completely liquid above the tipping temperature with solid precipitating below that temperature. For most of the work reported here, the 1100°C isotherm of the ternary was used. It was generally necessary to heat the solution to 50" to l00. C above the tipping temperature to dissolve all of the Gap in a reasonable length of time. The epitaxially grown layers were used both for optical transmission measurements to aid in the estimation of the way in which the absorption edge changed with aluminum concentration, and for the electron beam microprobe analyses to provide data for the determination of the solid solubility isotherm. RESULTS AND DISCUSSION Liquidus Isotherms in the Ga-A1-P Ternary Phase Diagram: Thermodynamic properties of the system. The only thermal effect studied in this work was that

Jan 1, 1970

By E. P. Whittemore

Experience with the ASC multipoint gas lift system was obtained in Colonia zone of the West Montalvo field near Oxnard, Calif. The wells in this pool produce from depths varying from 10,500 to 12,000 ft. Oil gravity is generally 14 to 15' API with a few extremes of 12 and 20" API. Some salt water is produced which causes some very viscous emulsions. Viscosities at 150F (which is the approximate wellhead temperature) vary from 5,000 to 100,000 SSU. Most of the production is by gas lift, although a few wells are produced by rod and hydraulic pump. About half of the gas-lift wells are on continuous flow and the remainder are on intermittent lift using large, ported, pilot-operated valves for single-point transfer of gas from casing to tubing. Gas-liquid ratios vary from about 6 to 10 Mcf/bbl of gross fluid lifted. Wells are produced to a 450-psi trap system. The following remarks will be confined to intermittent lift only, since this is the type of lift which has been achieved with the ASC valve system. The maximum gross fluid which has been produced by single-point intermittent lift is about 350 B/D in 3-in. tubing and 200 B/D in 21/2-in. tubing with gas-liquid ratios of approximately 7 to 9 Mcf/bbl. Some design changes could reduce this ratio. The ASC multipoint system has provided production as high as 480 BOPD in 21/2-in. tubing with gas-liquid ratios just under 4 Mcf/bbl. To be able to apply the multipoint system, it is recommended that a detailed explanation be obtained concerning transition-point pressure and stabilizer setting—what its significance is to the string design, how it may work for or against the operation of the well, how it is related to tubing sensitivity and how it affects the unloading operation. The unloading operation may only be of academic interest in a technical paper, but to the production foreman, unloading and setting the valves in operation is a very real problem and should be understood in detail. One item touched lightly in the paper was the unloading valve. This valve controls the maximum pressure at which the well can be operated. When lifting heavy viscous fluids, it is most important to set this valve for the maximum possible realistic operating pressure at the surface. If the well lifts easily, it is simple to set the ASC valves at a lower operating pressure and the unloading valve will remain closed; but if the well happens to be heavier to lift than anticipated, it may be desirable to operate on the unloading valve itself and use all the energy obtainable at the bottom of the hole. In the Colonia pool very heavy wet-gas gradients are experienced due to the viscosity of the liquid and the dense mist which is left behind a slug of fluid. There are many combination strings of 3- and 21/2-in. tubing. This aggravates the wet-gas gradient problem and provides wet-gas gradients of about 50 to 70 psi/1,000. An advantage which multipoint lift has provided is increased slug efficiency through better maintenance of pressure under the slug and decreased fall back as the slug passes up the tubing. By using multipoint injection, wet-gas gradients have been reduced to about 30 psi/1,000. This has reduced bottom-hole operating pressure and given a production increase. The ASC valve is not a simple device. It's operation is difficult to understand, and it must be understood to be used efficiently in gas-lift design. Operating problems are difficult to diagnose—whether they be caused by the fluid lifted, valve malfunction, lift gas rate, or operating pressure. Calculations and reasoning are required to find out what is causing the problem. Inherent in the ASC valve is the inability to create large pressure differentials across a slug. Large differentials may be required to overcome the inertia of very viscous fluid as it is being accelerated in the bottom of the hole. This is tied back to the design of the unloading valve and is one reason for the importance of setting the unloading valve for the highest possible operating pressure. ~u; to the narrow spread the ASC valves provide, it is impossible to cycle slower than about 24 cycles/day on choke control. If small production of 150 BOPD and less is expected, a surface time-cycle controller will be required if the most economical operation is to be achieved. To achieve a satisfactory operation, the operator must keep a record of any changes made in the operating pressure of the ASC valves. Anything which may cause changes in casing pressure in excess of the stabilizer setting will change the valve operating pressure, and if this is not noted from daily inspection of the well casing-tubing pressure recorder charts, the operator will lose control of the well. Significant results can be achieved using ASC valves; however, considerable knowledge is required to design with them, and attention to detail is required for satisfactory field operation.

Jan 1, 1965

By R. J. Fruehan, F. D. Richardson

Electrochemical measurements have been made of the activity of oxygen in copper and its alloys with silver and tin at 1100" and 1200°C. The galvanic cell used was Pt, Ni + NiO/solid ellectrolyte/[O] in metal, cermet, Pt The results do not support any of the equations so far designed for predicting the activities of dilute solutes in ternary solutions from activities in the corresponding binaries. If, however, a quasichemical equation is used with the coordination number set to unity, agreement between observed and calculated activities shows that this empirical relationship can be useful over a fair range of conditions. SEVERAL solution models have been proposed for predicting the activity coefficients of dilute solutes in ternary alloys from a knowledge of the three binary systems involved. Alcock and Richardson1 have shown that a regular model, and a quasichemical model,' in which the dissolved oxygen is coordinated with eight or so metal atoms, can reasonably predict the behavior of both metal and nonmetal solutes when the heats of solution of the solute in the separate solvent metals are similar. But when this is not so, neither model gives useful predictions unless coordination numbers of one or two are assumed. Wada and Saito3 subsequently adopted a similar model to derive the interaction energies for two dilute solutes in a solvent metal. Belton and Tankins4 Rave proposed both regular and quasichemical type models in which the oxygen is bound into molecular species, such as NiO and CuO in mixtures of Cu + Ni + 0. However, their models have only been tested on systems in which the excess free energies of solution of the solute in the two separate metals differ by a few kilocalories. Ope of the reasons why more advanced models have not been proposed is because few complete sets of data exist for ternary systems in which the solute behaves very differently in the two separate metals. For this reason measurements have been made of the activities of oxygen dissolved in Cu + Ag and Cu + Sn. Measurements on both systems were made by means of the electrochemical cell, Pt, Ni + NiO/solid electrolyte/O(in alloy), cermet,Pt [1] The activity of oxygen was calculated from the electromotive force and the standard free energy of formation of NiO, which is accurately known.5 Before investigating the alloys, studies were made of oxygen in copper to test the reliability of the cell and to check the thermodynamics of the system. Of the previous studies those by Sano and Sakao,6 Tom-linson and Young,7 and Tankins et al.8,7 have been made with gas-metal equilibrium techniques; those by Diaz and Richardson,9 Osterwald,10 wilder," Plusch-kell and Engell,12 Rickert and wagner,13 and Fischer and Ackermann14 have been made by electrochemical methods. EXPERIMENTAL The apparatus employed was the same as described previously,9 apart from slight modification. The molten sample of approximately 40 g was held in a high grade alumina crucible 1.2 in. OD and 1.6 in. long. The solid electrolytes were ZrO2 + 7½ wt pct CaO and ZrO2 + 15 wt pct CaO; the tubes 4 in. OD and 6 in. long were supplied by the Zirconia Corp. of America. They were closed (flat) at one end. In one experiment with Cu + O, both electrolytes were used in the cell at the same time. The reference electrodes inside the electrolyte tubes consisted of a mixture of Ni + NiO. They were made by mixing the powdered materials and pressing them manually into the ends of the tubes, with a platinum lead embedded in them. The tubes were then sintered overnight in the electromotive force apparatus at 1100°C. By sintering the powders inside the tubes (instead of using a presintered pellet9) better contacts were obtained between the electrolyte, the powder, and the platinum lead. Troubles arising from polarization9 were thus much reduced. The electromotive force was measured by a Vibron Electrometer with an input impedence of 1017 ohm; the temperature was measured with a Pt:13 pct Rh + Pt thermocouple protected by an alumina sheath. The couple was calibrated against the melting point of copper. The cermet conducting lead of Cr + 28 pct Al2O3, previously found to be satisfactory9 for use with Cu + 0 was also found satisfactory with Cu + Ag + 0 and Cu + Sn + 0. Superficial oxidation was observed, but it did not interfere with the working of the cell. The reaction tube containing the cell was closed at each end with cooled brass heads and suspended in a platinum resistance furnace. The tube was electrically shielded by a Kanthal A-1 ribbon which was wound round it, and the ribbon was protected by a N2 atmosphere between the furnace and the reaction tube. The cell was protected by a stream of high purity argon which was dried and passed through copper gauze at 450°C and titanium chips at 900°C. All the metals used were of spectrographic standard. Procedure. In studies of the system Cu + 0, be-

Jan 1, 1970

By A. Simkovich, R. W. Lindsay

The possibility of using sodium sulfide slags to remove copper from ferrous alloys has been investigated by Jordan1 and by Langenberg.2, 3 In these studies, such slags were determined to be capable of removing copper and sulfur from the melt. The present work represents additional effort to clarify the effects of temperature on copper removal. The experiments were performed in a 17-lb induction furnace. Graphite crucibles contained the melts and kept the baths saturated with carbon. Temperatures were measured with a calibrated optical pyrometer and were controlled by manipulation of power input to the furnace. Estimated accuracy of temperatures in this investigation is ± 10°C (18°F) for measurements prior to slag additions, and + 20°C (36°F) after slag formation. The procedure consisted of melting 800 g of electrolytic iron. During this step, powdered graphite covered the exposed iron surface. After a predetermined temperature was reached, copper shot was added. A sample of the molten alloy for chemical analysis was then aspirated into a silica sheath. Next, a slag-forming mixture of sodium sulfite and graphite was added instantaneously to the melt. The sodium sulfite amounted to one-tenth the charge weight of iron; sufficient graphite was added to combine with oxygen in the sodium sulfite, assuming formation of carbon monoxide and reduction of the sulfite to sulfide. Subsequent to the slag addition, the molten alloy was sampled periodically, with the exception of heat A in which no intervening samples were taken between the slag addition and the end of the run. The iron was poured into a graphite mold, and the ingots sectioned and drilled for samples. Results of selected heats are presented in Table I. Analyses of samples drawn from the iron prior to slag addition are listed under zero time. Two samples from heat D were reported with copper contents greater than the initial concentration in the bath. Owing to the gradual but complete disappearance of slag during this heat, it is believed copper momentarily became more concentrated in the upper portion of the bath while reverting from the slag. This is the region from which samples were drawn. It should be noted that analysis of the ingot was equal to the copper content at the time of slag addition. The terminal temperatures of heats D and E, and the initial sulfur content of heat A are also to be noted. Because of the large temperature drop which occurred when slag was formed in heat D, power input to the furnace was increased in heat E after the slag addition, causing a higher terminal temperature. In heat A, the initial sulfur concentration was relatively high as compared to heats B through E owing to contamination by some slag remaining in the crucible from a previous heat. It is evident from Table I that copper was removed at the onset of slag formation. Roughly 30 pct of the copper was taken into the slag, with the exception of heat D, which had approximately 50 pct removed. For a comparatively short time of slag-metal contact, it appears that no gain is to be made in copper removal through use of high or low temperatures. If the slag initially formed remains in contact with the iron for an extended period, temperature has a marked effect upon copper removal, as can be seen by studying results for the two extremes in temperature. At about 1425°C, the copper level remained relatively constant after the initial removal by the slag. However, in the region of 1670°C, a definite reversion of copper occurred. Reversion was incomplete in heat D, and complete in heat E. The final temperatures of heats D and E differed by about 75°C. This temperature difference is thought to be the reason for only partial copper reversion in heat D. It is believed the effects of temperature noted above are related to the evolution of a white fume, which appeared in every run except heat A. (In the case of heat A, the fume was practically indiscernible.) After each slag addition, a yellow flame formed for about 5 sec. When the flame subsided, a white fume appeared. Upon contact with surrounding cooler surfaces, this fume deposited as a white solid. In the experiments made at 1425°C, evolution of fume continued unchanged to the end of the runs. However, heats D and E exhibited a different behavior. A very noticeable decrease in fume evolution from heat D was observed. Furthermore, this heat had much less slag remaining than did runs A through C when the experiments were terminated. No slag remained at the end of heat E; evolution of fume from this heat ceased prior to pouring. Spec-trographic analysis of the white deposit indicated sodium to be the major metallic element, with the maximum concentration of iron and copper as 0.1 and 0.01 pct, respectively. It is supposed the white fume observed in these experiments is principally sodium oxide (Na2O), formed by oxidation of sodium in the slag and subsequent sublimation. (Sodium oxide is a white to gray substance in the solid state; at 1275oC, it sublimes.4) According to this mechanism, elevated temperatures would accelerate removal of sodium from the slag, sulfur pickup by the

Jan 1, 1961

By A. Totani, S. Nakajima, H. Okazaki

The phase diagram for the SnTe-SnSe system has been studied in the temperature range from 300° to 900°C by differential thermal and quenching techniques. The X-ray measurements were made on quenched specimens. High-temperature diffraction was also made to study the phase transition in SnSe. The system is proved to be of a eutectic type in which no intermetallic compound exists. The eutectic point is at the composition SnTeo.55 Seo.45. the eutectic temperature being 755°C. Solid solubility limits are SnTeo.6Seo.r and SnT eo. 3s Seo.6s at the eutectic temperature, and change almost linearly to SnTeo.aaSeo.lz and SnTeo.18 Seo.az as temperature decreases to 300°C. It was shown that the SnSe phase has a phase transition of the second order at about 540°C and that the transition temperature decreases with increase of the SnTe content. THERMOELECTRIC properties of tin telluride (SnTe) and tin selenide (SnSe) have been studied extensively in recent years. The variation of physical properties with composition could be of interest if these compounds form an appreciable crystalline solution. The purpose of present investigation is to confirm the formation of crystalline solution or intermetallic compound, if any, and to establish the phase diagram for this system. The crystal structure of SnTe is NaCl type with a cubic unit cell1 (a = 6.313A). The crystal of SnSe having an orthorh2mbic unit cellz (a = 11.496, b = 4.1510, and c = 4.4437A) is isomorphous with tin sulfide (SnS) which has a distorted sodium chloride structure. It has been known that SnSe has a phase at at 540°C; the transition has been assumed to be of the second order. As far as we know, only two studies on the SnTe-SnSe pseudobinary system have been reported. The conclusion obtained in these papers is that, in the composition regions near SnTe and SnSe, the system forms a crystalline solution of the SnTe structure and the SnSe structure, respectively, and that, in the intermediate region, both phases coexist. However, neither the variation of the solid solubility vs the temperature nor the liquidus and solidus were investigated. Hence present writers have attempted to determine the phase diagram of the system by differential thermal analysis (D.T.A.) and X-ray diffraction. EXPERIMENTAL Sample Preparation. Starting materials, SnTe and SnSe, were prepared by the direct fusion of commercially available high-purity (99.999 pct) elements. Stoichiometric amounts of each couple Sn-Te or Sn-Se were weighed into a clear fused silica ampule. After evacuation to a pressure below 10-3 mm Hg, the am- pule was sealed, and annealed at 900°C for 5 hr. The melt was quenched in water. X-ray analysis confirmed the formation of a single phase of SnTe or SnSe. The other samples, SnTel-,Sex were synthesized from these SnTe and SnSe by mixing them in the required ratio, followed by annealing at 900°C and quenching. These samples were used directly for D.T.A. For X-ray measurements, samples were annealed at 700°, 600°, or 500°C for 100 hr or at 300°C for 150 hr, and then quenched in water. It was found that the lattice constants of the SnTe phase annealed for 150 hr at temperatures above 500°C did not differ from those annealed for 100 hr at the same temperatures. However the X-ray phase analysis showed that at 300°C the annealing for 150 hr was necessary to attain a true equilibrium state. D.T.A. The solid-liquid equilibrium temperature was determined from D.T.A. measurements. The sample was sealed in an evacuated silica tube and molybdenum powders sealed in an another tube were used as a reference material. The sample and the reference tube were placed in a nickel block and were heated from room temperature to 900°C at a rate of 3°C per min and then cooled down at the same rate to 600°C. Thermocouples for these measurements were Pt-Pt. Rh (10 pct) and the error of temperature measurements was within + l0C. D.T.A. curves were obtained on a two-pen recorder and an automatic controller (PID type) was used for the program of heating and cooling. When temperature reaches the solidus from the low-temperature side, there appears an endothermic peak. The solidus temperature was determined by extrapolation of the straight portion of the starting flank of this peak to the base line. In a similar way, the liquidus temperature was determined from an exothermic peak on D.T.A. cooling curve. In the case of supercooling, if any, its degree can be estimated from the magnitude of the abrupt temperature rise. X-Ray . X-ray powder patterns were taken by a diffractometer using CuK, radiation. Since the SnSe crystal is cleaved easily, the powders become flaky when SnSe-rich samples are ground in an agate

Jan 1, 1969

By S. R. Goodman, Hsun Hu

A detailed study of texture developmenf in poly crystalline copper atzd 70-30 brass has been completed. Textural changes as a function of deformation are shoum by pole jigmres and by intensity measurements oF- various rejlectiotzs from the rolling plane and the rolling direction. These examinations were accompanied by measurements of stacking fault frequency, hardness changes, and microstructure. Some of the results were briefly presented earlier. Additional results reported here are consistent with the idea that deformation faulting or slip by partial dislocations is of primary importance in the formation of deformation textures in fcc metals. lo examine the idea that deformation faulting is of primary importance in determining whether the texture is the copper type or the brass type an extensive study of the development of polycrystalline textures in copper and 70-30 brass was initiated. Besides the determination of complete pole figures, the intensities of the various reflections from both the rolling plane and the plane perpendicular to the rolling direction, the peak shifts due to deformation stacking faults, and the hardness of the rolled specimens were examined at various reductions from 10 to 99 or 99.5 pct. Mi-crostructures were examined by transmission electron microscopy. Some of the results were briefly presented in an earlier publication.' Since then, additional information has been obtained. This is given in the present paper. EXPERIMENTAL PROCEDURE Material and Specimen Preparation. The material used was a commercial electrolytic copper bar 1i in. wide and 2 in. thick and a 70-30 brass bar la in. wide and 1i in. thick. Chemical analysis indicated a purity of 99.97 pct for the copper, with 0.025 pct 0 as the major impurity. The 70-30 brass was of higher purity with 0.0016 pct 0 as the major impurity. Extreme care was taken in the preparation of the starting material to insure uniformly fine grains with a nearly random initial texture. The two bars were first cold-forged and then annealed to eliminate any original texture. The grains were then refined by several cold rolling (approx 30 pct reduction) and annealing treatments. The + -hr anneals were carried out in a salt bath at 390" to 440°C for copper and at 490°C for brass. The resulting penultimate grain size was 0.06 mm for copper and 0.03 mm for brass, and both showed very little preferred orientation. The number of prior cold rolling and annealing cycles was such that the final thickness after various final reductions of 10 to 95 (for brass) or 99 (for copper) pct was the same (0.020 in.). These annealed strips were rolled in two directions by reversing end for end between passes according to the following schedule: 0.006 in. per pass to 0.100 in., 0.003 in. per pass to 0.050 in., 0.002 in. per pass to 0.025 in., 0.001 in. per pass to 0.020 in. Texture Determination. The development of rolling textures was studied by examining complete pole figures determined from the (111) reflection. Specimens thinned from one face of the strip to half thickness (0.010 in.) were used to obtain the central portion of the pole figures, while specimens thinned from both faces to 0.003 in. were used to obtain the peripheral portion. The reflection and transmission techniques have been described previously. In addition to X-rav pole figures, texture development was also studied b; examining the intensity variation of the (Ill), (200), (2201, (311), (331), (420), and (442) reflections from the rolling plane and from the plane normal to the rolling direction, as a function of deformation. The same specimens used for the central portion of the pole figures were used for the intensity measurements of the various reflections from the rolling plane. For intensity measurements from the plane normal to the rolling direction, composite specimens were prepared by mounting sections cut parallel to the transverse direction of the strip. An epoxy resin was used to bond these sections together, and the entire composite was then mounted in a cold-setting resin to facilitate subsequent polishing and etching to remove distorted metal at the cut. The intensities were expressed in units of the integrated intensities measured from an annealed copper specimen having almost no preferred orientation. Stacking Fault Frequency Determination. Following the analysis of Warren: the stacking fault frequency, a, was determined from the change in the peak separation (A%) of two neighboring reflections of a deformed specimen, as compared with the normal peak separations of a fully annealed specimen. To obtain sufficient intensities for the second-order reflections, (222) and (400), composite specimens were prepared from parallel sections cut from the strip at 30 deg to the rolling direction for copper and 25 deg for the brass.* From texture data, these sections are known to contain a large population of both (111) and (200) planes. Since residual stresses can also cause X-ray line shifts (the direction of line shifts depends upon the sign of the stress), the use of composite specimens consisting of sectioned planes should help compensate for these effects as the residual stresses change sign from the surface to the central section of a rolled strip. Since the amount of peak shift is almost un-measurable in brass rolled 15 pct and in copper rolled

Jan 1, 1969

By George M. Schwartz, Ian L. Reid

THE Soudan mine in the Vermilion district of northeastern Minnesota is the oldest iron mine in the state. It has shipped ore every year since 1884 and still contributes a yearly quota of high grade lump ore. No comprehensive report on the Vermilion iron-bearing district has appeared since Clements' monograph,' but Gruner2 discussed the possible origin of the ores in 1926, 1930, and 1932, and recently Reid and Hustad have added data on mining and geology .3, 4 For many years geologists of the Oliver Iron Mining Div., U. S. Steel Corp., have kept up to date a series of plans and vertical sections of the Soudan mine. In connection with mine operation considerable diamond drilling has been done, and this, together with the mine openings, has permitted a reasonably accurate picture of the structure of the orebodies and wall rocks. It has long been evident to geologists familiar with the mine that the ores were not a result of weathering, a point emphasized by Gruner in 1926 and 1930. As the deeper orebodies were developed it also became clear that replacement had played an important part in their development. In recent years it has been recognized that other iron ores were formed by replacement, as Roberts and Bartly5 have argued strongly for the deposits at Steep Rock Lake. On the basis of these facts G. M. Schwartz suggested to members of the Oliver staff that it would be desirable to study the evidence of replacement, particularly the possible alteration of the wall rock which would be expected if the replacement was a result of hypogene solutions. Rock Formations: The formations directly involved in the iron orebodies of the Soudan mine are few though far from simple. The country rock is largely the Ely greenstone of Keewatin age consisting of a mass of metamorphosed lava flows, tuffs, and intrusives which have been more or less altered by hydrothermal solutions. The predominant rock is chlorite schist. Interbedded with the original flows and tuffs are a series of beds and lenses of jasper to which the name Soudan formation has been applied. In the Vermilion district the term jaspilite has been used for interbanded jasper and hematite. According to modern usage these jasper or jaspilite beds do not comprise a formation separate from the Ely greenstone, inasmuch as the beds of jasper are interbedded with the flows and tuffs of the upper part of the greenstone. It would more nearly accord with modern usage to consider the Soudan beds a member of the upper part of the Ely formation. Because of incomplete rock exposure and exploration the number of interbedded jaspilite beds is unknown. In the mine, however, as many as nine major beds of jasper are known on a cross-section of one limb of the syncline, with an equal number on the other limb. In addition diamond drill cores show beds of greenstone down to half an inch in thickness. The thin beds are probably always tuffs. Structure: Rock structure in the Soudan area is complex, and because there are no recognizable horizons within the greenstone it is extremely difficult to work out the details. Generally speaking, the major regional structure is an anticlinorium, the axis trending east-west, with a westerly pitch. The Soudan mine is related to a synclinal structure on the north limb of the anticline about a mile from the west nose of the folded iron formation. The general structure at the mine is that of a closely folded minor syncline on the major regional anticline. A cross fault has dropped the east side so that the bottom of the syncline has not been reached, whereas to the west it is well shown by the mine openings and diamond drill exploration. Throughout the mine the beds of jasper, and ore-bodies that have replaced the jasper, normally dip northward at angles of 80" or steeper. In detail the jasper beds are extremely folded, probably as a result of deformation while they were still relatively unconsolidated. Orebodies: Ore in the Soudan mine is mainly a hard, dense, bluish hematite. Locally ore has been brecciated and cemented by quartz. The vugs commonly occurring near the borders of orebodies are lined with quartz crystals. They seem to have formed as part of the ore-forming process and are evidence that no folding or compression of the ore has taken place. The orebodies are numerous, varying greatly in size. Many lenses of high grade hematite are too small to be mined. Some of the larger orebodies have been followed vertically for as much as 2500 ft and horizontally up to 1500 ft. The large ore-bodies are extremely irregular in outline in the plane of the beds of jaspilite. In width they are more regular, as they are strictly governed by the width of the jaspilite beds and the greenstone wall rock, which seems to have resisted replacement by hematite. At many places the orebodies replace the jaspilite completely and have a footwall and hanging wall of greenstone. At other places either one or both walls may be jaspilite. Geologists who have studied the orebodies in recent years agree that evidence for the replacement origin of the hematite bodies seems conclusive. AS noted above, many of the orebodies replace jaspilite beds from wall to wall with no evidence whatever of compaction. The replacement origin is also supported by details of the banding which is characteristic of the

Jan 1, 1956

By W. Bonfield

The temperature dependence of the microscopic yield stress (the stress to produce a plastic strain of 2 x 10-6 in. per in.) and the stress-plastic strain curve of polycrystalline Cu 1.9 wt pct Be have been measured for the solution treated condition, an intermediate condition containing G.P. zones and ?' precipitate and the overaged ? precipitate condition, in the range from -58° to 200° C. A transition in micro -yield behavior and a large temperature dependence were noted for the intermediate condition, which are interpreted in terms of the interaction of glide dislocations with two differently sized zones. In comparison the microscopic yield stresses of the solution treated and overaged conditions were less sensitive to temperature variations and are satisfied by the Mott-Nabarro and dislocation bowing theories, respectively. A determination of the temperature dependence of the yield stress of a precipitation hardening alloy has provided a powerful tool for evaluation of the operative deformation mechanism. There is a marked contrast between the effect of temperature on the yield behavior of a metal containing coherent zones or intermediate precipitates, which can be "cut through" by mobile dislocations, and a metal containing a dispersion of noncoherent particles, through which dislocation "bowing out" is the dominant role of deformation.' These studies have in general been confined to single crystals, as it was considered that similar experiments on polycrystalline material did not produce good data because of the lack of sensitivity with which the yield stress could be determined. However, this objection has been removed by the introduction of mi-crostrain techniques, with which the yield stress in polycrystalline materials can be measured to a strain sensitivity of 10-6. Such measurements have not only shown that the deformation of polycrystalline precipitation hardening alloys can be examined with the same detail as single crystals, but also that some unexpected results are obtained.' In this paper the results obtained from a study of the temperature dependence of the microscopic yield stress (the stress to produce a plastic strain of 2 x 10-6 in. per in.) and the stress-plastic strain curve of a polycrystalline Cu 1.9 wt pct Be precipitation hardening alloy (Berylco 25) are discussed. The temperature dependence of the alloy was measured for three different conditions: 1) The solution treated condition (a supersaturated solid solution of a containing ~12 at. pct Be3) which is obtained by water quenching the alloy from 800° C. 2) The condition of y' intermediate precipitate, to- gether with some G.P. zones,' which is produced after an aging treatment of 2 hr at 315°C from the solution treated condition. (The alloy was cold rolled to 40 pct reduction prior to aging to minimize grain boundary precipitation effects.)4 3) The condition with equilibrium ? precipitate structure2 which is developed after an aging treatment of 24 hr at 425° C. EXPERIMENTAL PROCEDURE Tensile specimens of gage length 1 in. and with rectangular cross section of 0.18 by 0.06 in. were prepared from the solution treated, cold rolled alloy and were either resolution treated for 1 hr at 800°C, followed by water quenching, or aged for 2 hr at 315°C and 24 hr at 425° C to produce the desired precipitate structures. The microstrain characteristics of the aged specimens were determined at temperatures from —58" to 200° C and those of the solution treated specimens from -58° to 30° C. Each temperature was controlled to ± 0.2°C, which was a level of stability sufficient to eliminate thermal expansion effects from the measurements (~1.2°C temperature increase produced an extension of 2 x 10-6 in.). The microplastic behavior of the specimens in the temperature range below 82" C was measured with a standard Tuckerman strain gage,5 while at temperatures above 82°C a modified Tuckerman gage with a reduced strain sensitivity (4 x10-6 in. per- in.) was used. A load-unload technique was used to establish values of the microscopic yield stress. The specimen was strained at a constant cross head speed of 2 x 10-2 in. per min to a given stress level, at which the total strain was measured. Then the specimen was immediately unloaded at the same rate and any residual plastic strain determined. This procedure was repeated for an increasing series of stress levels until the microscopic yield stress was established by a direct measure of the stress to produce a residual plastic strain of 2 x 10-6 in. per in. (It should be noted that, as reversible dislocation motion occurs at stresses less than the microscopic yield stress,2 the plastic strain rate at this level was not constant.) In an ideal test, the microscopic yield stress would be determined from a continuous stress-strain measurement, rather than from a load-unload sequence, in order to eliminate mechanical recovery effects.6 However, it was found experimentally that mechanical recovery was negligible in Cu 1.9 wt pct Be at small plastic strains for all the temperatures investigated, as the microscopic yield stress was independent of the number of load-unload cycles employed (i.e., the values measured for specimens subjected to different numbers of cycles was within the experimental scatter determined for specimens tested in an identical manner). Therefore, it is reasonable to consider the microscopic yield stress determined in the load-unload

Jan 1, 1969

By A. L. Baxley

An Untapped Energy Resource As much as 20 billion cubic feet of natural gas per day are flared from remote oil fields for lack of a commercially viable means of capturing, transporting, and marketing such gas. The magnitude of these gas flares can be put into perspective from an early satellite photograph (Fig. 1) which shows lights from the major cities of Russia and Eastern Europe dwarfed by the natural gas being flared in the Persian Gulf. Together, these wasted resources contain the energy equivalent of about one-half of the gasoline consumption in the United States today (Fig. 2). Additional trillions of cubic feet of natural gas are "shut-in" because of no economically viable means of commercial recovery. Methanol and liquified natural gas (LNG) are the only two practical fuel products which can be produced economically from these gas supplies. Many of these gas supplies are less than 500 million cubic feet of gas per day, making an LNG facility uneconomic. In contrast, barge-mounted methanol plants can economically convert billions of cubic feet of gas per day into safe, clean-burning methanol. The methanol approach offers the only economical route to transform vast, known reserves of natural gas into a highly versatile primary liquid fuel. Methanol Barges: An Innovative Solution The barges will be towed to suitable offshore and upriver locations such as Alaska, South America, Africa, Southeast Asia, Australia, New Zealand, and the South Pacific Islands, as well as fields in the Persian Gulf and Mediterranean Sea. At the offshore production site, a barge will be anchored by a single point mooring buoy that will also serve as an entry point for natural gas feedstock and an offloading point for methanol (Fig. 3). At some sites the barge would be beached. Each barge will produce methanol and store it in internal tanks with a capacity of 18 million gallons. The methanol will be offloaded into conventional tankers and safely transported directly to market. Unlike LNG, methanol requires neither specially built carriers nor specially built receiving terminals. Once a particular gas field has been exhausted, the barge will be towed to another location to continue production. Each barge will measure 320 by 500 ft, approximately the size of four football fields, and will have the capacity to produce 1 million gallons or 2800 metric tons of methanol per day, from approximately 100 million cubic feet of natural gas per day (Fig. 4). The barges will use the highly successful "low- pressure" design developed by the Lurgi Company of Germany, a process proven in land-based methanol plants throughout the world during the last ten years. The decision to use Lurgi technology for "sea-trans- portable" methanol plants was based on the higher efficiency and greater operability of the technology compared to other commercially proven processes. The conversion plant will be designed to accept a wide variety of feed gas compositions, and will produce chemical-grade methanol for the broadest market base (Fig. 5). To minimize costs and construction time, the barge-mounted plants will be built in the high technology environment of a domestic or foreign shipyard. Selection of the construction site for each barge will be dictated by the location of the production site and by the relative construction costs. A number of shipyards have the capacity to build several barges per year. The detailed marine engineering to integrate the design of the processing plant with the floating platform can be performed by numerous major engineering companies around the world. Production Economics The barge-mounted plant concept not only assures large volumes of methanol, but it also keeps the overall production cost low by minimizing construction cost and providing access to low cost natural gas feedstock with no alternative or a negative value. Together, these advantages make the barge-mounted methanol plants economical today. The cost structure of a new barge-mounted methanol plant differs from that of existing methanol producers around the world (Fig. 6). For example, if a U.S. Gulf Coast producer is paying $4.70/MMBtu in 1985 for natural gas, the barge plant could afford to pay about $1.6O/MMBtu for gas and be able to deliver methanol to the Gulf Coast at the same price. At some future date such as 1990, a gas cost of $6.70/MMBtu for a domestic producer would have cost parity with about $3.60/MMBtu gas cost for the barge plant. In many foreign markets, feedstocks other than natural gas are used for methanol production (Fig. 7). For example, most of Japan's capacity is based on LNG while Western Europe uses residual oil or naptha. Because these feedstocks are substantially more ex- pensive than natural gas used by U.S. producers, the barge plants will compete even more favorably in these foreign markets. As crude oil prices rise, the value of methanol in each of these markets will increase. However, the hierarchy of methanol values in these markets should remain unchanged. Furthermore, the cost advantage for using methanol in these markets will improve as world energy costs increase since the value of remote gas should not escalate significantly.

Jan 1, 1982

By O. J. Hassel, W. T. Maidens, J. H. Calbeck

The rotary gas fired reheating furnace used by the American Zinc Oxide Co. at Columbus, Ohio for Therotarygasfiredreheatingfurnacerefining lead-free zinc oxide is described. The outstanding features of this operation are that the color of the zinc oxide is greatly improved, sulphur is eliminated, and cadmium arethatrecovered without densifying the product to an objectionable degree. IN 1919 Leland S. Wemple obtained a patent for a process of reheating zinc oxide wherein the "coarsening of grain due to excessive heating was avoided." He taught in his specification that if solid carbonaceous material, such as lamp black, was added to the zinc oxide in proper amounts prior to reheating, objectionable sulphur compounds could be removed and the color would accordingly be improved and no objectionable densification would occur because of the relatively low temperature required. The situation that made this invention imperative was the newly opened zinc oxide plant of the American Zinc, Lead & Smelting Co. in Hills-boro, Ill. This was one of the early Western Type American Process zinc oxide operations. Characteristic of all of these early Western operations using Tri-State and Western ores was the great difficulty encountered in obtaining a product low enough in sulphur to compete with the Eastern Type American Process zinc oxides which were made from ores containing very low sulphur percentages. Wemple demonstrated that the refining process of his invention produced a superior color and although this was true and a most welcome feature, the primary purpose of the early refining operations at Hillsboro was to reduce substantially the high sulphur content of the crude zinc oxide. Although many and varied attempts had been made for refining zinc oxide none of the processes had a commercial history of any consequence until Wemple's invention became standard practice for the American Zinc, Lead & Smelting Co. in 1919 and their operations have been unique in that substantially all of their lead-free zinc oxide has been reheated since the first installation at Hillsboro. This process has become known in the industry as refining. The furnace developed by Wemple and continued in use by the company from 1919 until 1943 was unusual and merits some consideration by way of review in this paper. The furnace was essentially a double hearth coal-fired muffle furnace with a mechanical raking system consisting of a central shaft supporting six rabble arms in each muffle. The untreated or "crude" zinc oxide was fed onto the outer rim of the top muffle, moved to the center where it dropped to the lower muffle and progressed to the outer rim where it was discharged into an alloy screw conveyor. The retention in this furnace was extremely short, about 5 min, and the shallow zinc oxide bed on the hearths of the muffles was being continuously turned by the fast moving rabbles. Soft coal was burned on the grates below the lower muffle and the long yellow flame necessary to carry the heat around both muffles resulted in a very inefficient combustion of the fuel. The temperature of the top of the lower muffle seldom exceeded 65 °C although the oxide itself often reached 700°C before discharge. The capacity of this furnace was approximately 1/2 ton per hr. In our plant at Columbus it was necessary to keep four of these furnaces running in parallel to take care of the production because, as mentioned above, every pound of zinc oxide produced during these 24 yr passed through one of these refining furnaces. An essential part of this refining operation was the use of carbonaceous material admixed with the zinc oxide fed to the furnaces. Between 1 and 2 pct of a bran produced in the processing of cotton seed was added to all zinc oxide charged to the furnaces. The bran ignited on the top hearth and was still burning when the charge fell from the top hearth to the bottom hearth making a cascade of sparks. The rapid turning of the zinc oxide caused these particles of bran to flash on the hearths behind each rabble; but the combustion, of necessity, had to be complete by the time the charge reached the outer rim of the bottom hearth, otherwise the finished product would be contaminated with the charred particles of bran which would give the zinc oxide an unsatisfactory color. Although this operation was initiated to reduce objectionable sulphur percentages, as time went on new properties of the product were appreciated which made advisable continuing the refining process long after other methods of sulphur reduction became known in the industry. The particle size and particle size distribution, the absence of colloidal fines and perhaps a unique surface condition gave this product an outstanding performance when used in paints. The Wemple furnaces installed in Columbus in 1919 had to be rebuilt frequently and were extravagant in the use of fuel. The raking mechanism and the muffles required excessive maintenance expense and as the furnaces wore out the problem arose whether to continue along this line or to explore the possibilities of obtaining similar or better results in the simpler and more commonly used rotary furnace. To this end special research was initiated in 1941 on a small laboratory rotary

Jan 1, 1951

By Frederick F. Frick

SPONGE iron as produced at Anaconda is a fine, -35 mesh, impure product, about 50 pct metallic iron, obtained from the reduction of iron calcine at a temperature of 1850°F by use of coke resulting from slack coal. The metallic iron particles are bulky and spongey and precipitate copper readily and rapidly from a copper sulphate solution. Investigation of the treatment of Greater Butte Project, Kelley, ore at Anaconda early showed the desirability of using sponge iron as a precipitant for the copper in solution resulting from desliming of the ore in a dilute sulphuric acid solution. Anaconda had done considerable work on the production of sponge iron in 1914 for use as a precipitant of copper from leach solutions. Some success and considerable experilence were attained at the time. indicating that, sponge iron might be successfully made by a modification of the process used in 1914, a batch process in which an iron calcine was reduced by means of soft coke, resulting from noncoking coal, in a Bruckner-type revolving horizontal cylindrical furnace widely used 50 years ago. The coke and calcide formed the bed in the Bruckner furnace, which was rotated at about 1 rpm. The bed was brought to a temperature of about 1800°F by means of an oil flame over the surface. Although results were reasonably satisfactory, they did not warrant full development of the process at that time. A good deal of work has been done in the last 50 years on the production of sponge iron. The objective in some cases has been the production of a precipitant for copper from solution, but the bulk of the work has been done for the production of open-hearth steel furnace stock. The production of an open-hearth stock presents two problems rather than one: first, producticon of the sponge iron, and second, what is perhaps of equal difficulty and importance, conversion of the sponge iron into a form suitable for use in the open-hearth furnace. So far as is known to the writer, none of the sponge iron processes tried in the past have proved to be economically feasible. However, Anaconda had a combination of conditions appearing to justify an attempt to produce sponge iron which would serve for the leach-precipitation-float process. It was thought that the process used in 1914, if changed to a continuous one, might work out satisfactorily. The following favorable conditions at Anaconda justified the investigation: 1—A sufficient tonnage of good grade iron calcine resulting from the roasting of a pyrite concentrate in one of the acid plants, at substantially no cost. 2—Reasonably cheap natural gas. 3-—The fact that there was no need for production of a high grade product. 4— The fact that there was no need for obtaining a consistently high reduction of' the iron in calcine. A small revolving Bruckner-type furnace about 2 ft ID by 4 ft long was set up for early pilot work at the research building. This pilot furnace showed that a satisfactory product could be obtained at reasonable cost. It also indicated a marked advantage in preceding the reduction furnace with a furnace of similar size and capacity for preheating and roasting out any residual sulphur from the feed. The small furnace was operated for several months, various details of the process were worked out. and sponge iron was produced to supply a pilot LPF plant which treated 300 lb of Kelley ore pel- hr. Later a second pilot furnace 5 ft in diam and 12 ft long inside was set up at our reverberatory furnace building. This furnace confirmed the data of the small furnace and gave a basis for design of the final plant. At Anaconda a pyrite concentrate, running about 48 pct S, is recovered from copper concentrator tailings by flotation. This concentrate is roasted to sulphur of 3 pct or less at the Chamber acid plant. The iron calcine contains about 57 pct Fe and 18 pct insoluble. The iron calcine feed, as mentioned before, is re-roasted and preheated in a reroast furnace preceding the reduction furnace. Both are of the Bruckner type. The reroasted calcine is fed into the reduction furnace at 800" to 1000°F along with 30 pct slack coal. In the feed end of the furnace the volatile is burned from the slack, giving a soft coke which readily serves for reduction of the iron. Hard metallurgical coke will not serve the purpose. since it does not reduce CO readily at a temperature of 1850°F. All indications are that the actual reduction of the iron is accomplished by carbon monoxide below the surface of the bed, which is 30 in. deep at its center. Apparently there is a constant interchange: Fe²O³ + 3CO = 2Fe - 3CO², CO² + C = 2CO Actually iron oxide is reduced by CO at somewhat lower temperature than the 1850 °F used in the process. but this temperature is necessary to obtain a satisfactory rate of furnace production. The furnace atmosphere is generally reducing, and typical blue carbon monoxide flames satisfactorily cover the bed. Gas flames from four 3-in. Denver Fire Clay Inspirator burners are played directly on the bed, which is slowly cascaded by the 1 rpm of the furnace. An excess of coke is necessary to assure maintenance of good reducing conditions in the furnace bed. Part of this coke is recovered for re-use.

Jan 1, 1954

By K. Prins

ONE of the more pressing problems faced by the coal industry today is the development of adequate means for meeting conservation laws, particularly those involving stream pollution, in various parts of the country. Discharge of dirty coal washery water into streams and rivers is almost universally frowned upon. Many states have enacted laws carrying heavy fines to curb the practice. The Prins stream-cleaner is one of the latest machines to enter the market. It is closely related to the cyclone thickener in principle. Eleven stream-cleaners are currently operating, ranging in size from 4 to 16 in. diam. In more recent installations the water enters directly in line and on a tangent with the impeller. The impeller consists of a vertical shaft up through a packing gland and bearings, and a V-belt pulley. The lower part is a tubing fastened to the shaft above, extending through the water intake compartment and provided with six vertical flat bars welded to the tubing. Portholes are situated in the upper end of the tubing, immediately below the point where shaft and tubing join together. The portholes are placed so that they are in open communication with the upper compartment of the stream-cleaner from which the processed water is discharged. The impeller is motor driven with a wide range vari-pitch drive employed between motor and impeller. The motor is mounted vertically, and the mounting provided with a vertical hinge allowing for needed adjustment of the wide range vari-pitch drive. The dirty coal washery water entering the machine under 20 lb psi pressure, flows from the compartment above the impeller, between the impeller blades, and is whirled around in the vertical section of the impeller enclosure. The speed of the impeller supplies centrifugal force and velocity required for separating suspended solids from water. The lower part of the machine consists of a cone, whose action is similar to other machines of the same type. The underflow discharge orifice is a cold rolled steel block machined to correspond with the cone angle and allows insertion of steel tubes of different diam. On 16 in. machines a 1 ¾ in. ID vertical discharge pipe is used. Provision is made for attaching a curved section of the same diam to the vertical pipe, to which, in turn, different lengths of horizontal pipe can be connected. Curved Pipe Advantageous It has been found that a curved pipe offers resistance to discharged material flow. In addition, the rotary motion of the underflow can be easily arrested in a curved pipe. Impeller speed of the 16 in. diam machine is provided from 400 to 800 rpm. A speed of about 474 rpm is suitable for maintenance of a constant underflow in coal slurry. In one installation 5x ¼ in. coal is cleaned in a Jeffrey Baum type washer at a 225 tons per hr rate. Washer installation is of the conventional type and a drag type sludge tank is used for water clarification purposes. Capacity of the water system, including the Baum washer, is about 40,000 gal. Before placing the stream-cleaner in operation, it was necessary to flush out the entire system every five days of two shift operations. The only time the system is drained now is for repair work on the sludge conveyor or the rig. The suction line of the stream-cleaner pump terminates in a number of small branch lines located at a depth of about 4 ft above the sludge conveyor. Each branch line extends the full width of the tank and is provided with four intake ports, each one with a funnel shaped inlet projecting downward. The arrangement provides an extensive pick-up area, for dirty water, and the inlets are arranged for a low rim velocity, preventing the taking in of coarse particles. The funnels are also arranged to extend up or down in the tank. They are set to pick up -60 mesh material exclusively.' The material is a high ash and high sulphur product and thus has to be disposed in the refuse conveyor. The underflow of the stream-cleaner is discharged on top of the washery refuse which is carried in a drag type horizontal conveyor, discharging into another refuse conveyor inclined at 30º with a short horizontal loading section. Some Disadvantages The impeller inherits certain disadvantages because of the nature of its construction. Additional moving parts make it subject to wear and maintenance costs. The advantage of being able to maintain a constant speed, however, to produce desired water velocity in the machine outweighs the drawbacks. Better separation between water and solids can be obtained by regulating time of residence of water through adjustment of valves in the intake and discharge lines. The amount of fines encountered during plant operation will vary because of higher or lower moisture in coal passing over fine coal vibrating screens. Even the amount of fines picked up by underground loading machines will be inconstant. Consequently, the percentage of solids will vary in water to be processed. The velocity in the feedlines to the slurry thickeners will fluctuate, with the required water velocity lacking. Another advantage advanced for the machine is its ability to operate on 15 to 25 lb line pressures at the water intake, reducing pump power required and pump maintenance.

Jan 1, 1952

By A. A. Collins

When we speak of high copper on a lead blast furnace we think in terms of 4 to 5 pct, or. any lead charge carrying over 1 pct. Any copper on charge will produce its corresponding troubles such as lead well, extra slag losses, drossing problems, and the working up of the dross. This is indeed a very interesting subject and one that used to give the old-time lead metallurgists such as Eiler, Hahn and lles many worries, not so much in the actual operation of the hlast furnace but in the working up of the copper. When the American nletallurgists commenced with the American rectangular-shaped lead blast furnace in the 1870's and got away from the reverberatories such as were in use in Germany and other parts of the world, they went to greater tonnages, as 80 to 100 tons per day in comparison to the 20 to 30 tons per day in the other processes. With the greater tonnages along with insuficient settling capacity, the silver losses in some cases were increased. Hence the lead-fall was low, for there were no leady concentrates in those days to assist the metallurgist to gain lead or an absorber for the precious metals; and in some cases copper sulphides were added intentionally to the charge to produce a copper matte to lessen the silver losses through the dump slag. The operators in those days thought that where some copper was always present in the lead ores the copper should not enter into the reduced lead and alloy with it. This, by the way, is just the reverse of our present-day practice, when we try to put all of the copper into the blast furnace lead and to remove the same through the drossing kettles. Therefore the furnace was operated to produce a certain amount of matte or artificial sulphides, since, due to the great affinity of copper for sulphur, any copper present would enter the matte almost completely. Thus, the lead bullion produced was practically free from copper. The products of the furnace were metallic lead or lead bullion, containing 05 to 95 pct of the lead and about 96 pct of the silver which were in the ore—a lead-copper-iron matte which contained nearly all the copper in the ore and the slag, the waste product. In the United States, up through the year 1092, we find the small furnace 100 X 32 1/2 in. with 12 tuyeres, some 6 on each side, plagued with a small amount of poorly roasted sulphides— either from heap or hand roasters that produced matte. This matte was roasted and if poor in copper was returned for the ore smelting. Otherwise it was smelted either alone or with additions of rich slags or argentiferous copper ores, the products being lead and a highly cupriferous matte, the latter being subsequently worked up for its copper. The lead metallurgists kept trying and improving on furnace and roasting equipment designs until we find blalvin W. Iles constructing at the old Globe Plant at Denver what came to be the modern furnace. That is, in 1900 he built a furnace of 42 in. width by 140 in. at the tuyeres with a 10 in. bosh and a 16-ft ore column. This type has been more or less standard to the present time, though modified in width and length to meet the demand for large tonnages and improvements in structural details. In 1905 at Cananea, Mexico, Dwight and Lloyd developed the present down-draft sinter machine that has meant so much in producing a well-processed material for the lead blast furnace. In 1912 Guy C. Riddell came forth with double roasting at the East Helena Plant of the American Smelting and Refining Co., which removed the "zinc mush plague." Incidentally, with the introduction of double roasting, which most lead plants were forced into after 1924, when lead flotation came into its own, less matte or no matte was produced. When this stage arrived, the copper was forced into the dross and the casting of lead at the blast furnace lead-wells was stopped. In plants with a fair copper carry 1 pct or better on the blast furnace charge, the lead wells became inoperative once the production of matte stopped. The copper drosses clogged the lead wells and even with bombing, either water or dynamite, the operators could not keep them open. Thus, the lead wells were abandoned in some plants, such as at the El Paso and Chihuahua smelters of the American Smelting and Refinillg Co., and all lead taken out through the first settlers. The elimination of sulphur, espccially sulphide sulphur, from the blast furnace charge and the nonproductiori of matte resulted in a great saving of tinie, energy and equipment in the recirculation of the copper, With the copper content in the dross and dross-fall ranging in quantities from a few percent up to 60 pct, such as at El Paso, a drossing problem was created. As the old-time operators hated dross and buried the same in the shipping bullion, the modern metallurgists from 1925 on decided that with increasing dross-falls they would have to adopt the lead refiner's ideas of drossing kettles with subsequent treatment of the lead with a sulphur addition to have the shipping lead of 0.01

Jan 1, 1950

By George F. Learning

TOTAL DIRECT IMPACT The uranium mining industry's principal economic impacts on the Texas economy are the result of three flows of money from the industry into the remainder of the state's economy. These three are: (1) money paid to individuals (personal income) ; (2) money paid to other businesses (business income); and (3) money paid to state and local governments (government revenues). As these direct payments from the uranium industry to various other sectors of the Texas economy subsequently circulate and recirculate within the state, the indirect effects of uranium mining's direct impacts multiply to reach amounts significantly higher than the direct income flows alone. Over the past decade, the uranium mining industry has substantially increased its role as a provider of jobs, personal income, business income, and government revenues in Texas. The growth has come almost exclusively in a largely rural, seven-county area that lies within the triangle formed by the Laredo, San Antonio, and Corpus Christi metropolitan areas. The uranium mining industry, in fact, has been the major dynamic element in this rural area despite relative stagnation in most of the region's other basic economic sectors. Over the three years from 1976 to 1978, the South Texas uranium mining industry directly contributed a total of $115 million to the economy of the seven- county region in which it operated and $164 million to the economy of the entire state of Texas. In 1979 alone, the total direct contribution of the industry to the Texas economy had climbed to $124 million in personal, business, and government income. PERSONAL INCOME IMPACT In the period from 1976 through 1978, the South Texas uranium mining industry provided an average of $12.5 million in personal income each year directly to residents of Atascosa, Bee, Duval , Karnes, Live Oak, McMullen, and Webb counties -- the seven Texas' counties that make up the South Texas Uranium Belt. A1 though 84 percent of this resulted from the employment of area residents in uranium industry jobs, some amounts were also provided by the payment of rents and royalties to land owners for the use of their land and mineral rights in uranium mining operations. In 1979, the uranium industry provided approximately $38 million to residents of the Uranium Belt and the rest of Texas. This was more than double the average of $16.1 million provided to Texas residents during the 1976 to 1978 period. The full importance of the uranium industry as a source of personal income, however, should not be reckoned merely by the amount of wages and salaries that it pays directly to its own employees, nor by its rent and royalty payments paid directly to land and mineral rights owners living in Texas. The added payments that the industry makes directly to other Texas businesses and state and local governments in Texas are themselves converted into personal income as those business firms and government units in turn pay their employees. All of the direct income payments made by the uranium industry circulate and recirculate within the state's economy, multiplying their impact as they go, until they eventually all leak out of the state as federal taxes or as payments to individuals or businesses located outside of Texas. The combined circulation and recirculation of the direct personal, business, and government income that was provided by the-uranium industry in Texas during 1976, 1977, and 1978 resulted in an average annual amount of indirect personal income of more than $83 mill ion. This alone was $20 mill ion more than the industry's average annual sales during the same years. The total of combined direct and indirect personal income contributed to the Texas economy by the uranium mining industry in that same three-year period thus averaged almost $100 million annually. In 1979, the amount of indirect personal income contributed to the Texas economy by the circulation of uranium mining's direct contributions had risen to about $196 million, more than double the average of the previous three years. The combined direct and in- direct personal income impact in 1979 thus amounted to $234 million. BUSINESS INCOME IMPACT The income provided directly to other Texas business firms through the purchase of needed goods and services by the uranium industry has been twice as big as the industry's payrolls. In 1976, 1977, and 1978, the South Texas uranium industry spent an average of almost $36 million each year to buy both goods and services from other Texas businesses. By 1979, this direct contribution to the Texas economy had swollen to $76 million. The biggest share of the uranium industry's payments to other businesses have gone to contracting firms, including both construction firms and those providing specialized mining services. In the past four years, about 40 percent of the direct payments made by Texas uranium producers to other Texas firms have been to contractors. Texas wholesale and retail firms have also shared in the business sales provided by the South Texas uranium mining industry. Over the past four years, Texas wholesale and retail trade businesses have accounted for about 34 percent of the uranium mining industry's purchases from other Texas businesses. Public utilities firms have received another six per- cent, while Texas manufacturers and transportation firms have accounted for about five percent each. The other sectors of the state's economy, including other mineral industries, agriculture, finance, insurance, real estate, and services, have accounted

Jan 1, 1980