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By J. B. Cheatham

Applications of the mathematical theory of plasticity promises to lead to the solution of many drilling and rock mechanics problems. ,Because of mathematical considerations, the inelastic behavior of rock has frequently been represented by a perfectly plastic model in conjunction with a yield criteria of the Coulomb or Mohr type. The totality of all stress states for which a solid ceases to behave elastically can be represented as a limit surface in stress space. ,Probing of such limit surfaces indicates details of strain hardening which are not provided by the standard triaxial testing procedure. Probing tests of the limit surfaces have been performed on Cardova Cream limestone to provide data for extending plasticity theory to cover situations in which consolidation and strain hardening are present. Test results indicate that this highly porous limestone undergoes a permanent volume decrease when it is subjected to hydrostatic pressures in excess of 3,500 psi. A virgin sample tested under a confining pressure of 1,500 psi has a yield strength of 1,700 psi; however, if the sample is subjected to a consolidation pressure of 5,000 psi, before testing at 1,500 psi, the yield strength is raised to 2,300 psi. Thus, both consolidation and strain hardening are important considerations in describing the mechanical behavior of this limestone. Tests conducted with the axis of the core having different orientations indicate that this rock is also anisotropic. Portions of the initial and subsequent limit surfaces are determined for samples loaded either perpendicular or parallel to the bedding planes. INTRODUCTION Previous experimental work in rock mechanics indicates that no mathematically tractable constitutive theory is inclusive enough to describe the mechanical behavior exhibited by all types of rocks under all conditions of stress and temperature. Indeed, the type of deformation encountered in a single type of rock is known to depend upon the stress and temperature conditions in the rock during deformation.l-5 Certain rocks and minerals, notably those minerals composed of ionic salts, have been shown to exhibit plastic deformation when tested under conditions of high confining pressure.2 Since the mathematical theory of plasticity provides simplifications over the theory of elasticity in certain types of problems, such as those in which limit analysis can be applied, it is of interest to know under what conditions plasticity theory may be applied to rock mechanics problems. The following factors determine the nature of the deformation a particular specimen will undergo: (1) the microscopic structure of the rock, i.e., the structure visible under an optical microscope, including number of phases, porosity, distribution of phases, (2) the mineralogical structure of the solid phases, (3) the conditions of stress and the rate of change of stress and (4) the temperature. Extensive experiments on Yule marble, Carthage marble, Solenhofen limestone and other calcerous rocks indicate that these relatively nonporous rocks deform plastically under certain conditions of loading, and creep under other conditions of loading.l-3 This study is concerned with the behavior of Cordova Cream limestone (Austin chalk) which is also composed almost entirely of calcite and thus has the same mineralogical composition but, because of a rather large porosity, it possesses a different microscopic structure. This investigation was undertaken to learn if Cordova Cream limestone deforms plastically despite the embrittling effect of the pore spaces, and to provide data which can be used to determine whether the mathematical theory of plasticity can describe the mechanical behavior of Cordova Cream limestone. A brief discussion is given of some of the underlying principles of plasticity theory concerning stress space and limit surfaces. Strain hardening is defined and a discussion is given of how the mechanical behavior of a material can depend on the previous history of the loading. This is followed by a description of the experimental procedures

By Robert Y. Grant

This study outlines the history of metal mining in Japan and the characteristics of the industry as they existed from 1925 to 1945. Mining and milling operations are described together with the role of the Japanese Government. A description of post-hostilities conditions and the present state of the industry is included. MANY of Japan's larger and better known mines date back to the early centuries of the Christian Era. The first stimulus to mining in Japan came through the growth of Buddhism, A. D. 600 to 700, and the attendant need for copper and other metals for temple ornaments and statues. During the Feudal period, 1333 to 1568, mine operations were expanded to meet the demand for ornament metal and for financing of campaigns. After removal of the seat of government from Kyoto to Tokyo, about 1600, the Tokugawa Shogunate placed great emphasis on mining, and many new mines were opened, some of which are still operating. Toward the end of the Tokugawa period, mining had declined sharply from its earlier peak. The easily worked, high grade portions of the orebodies had been exhausted, and the mining techniques then available could not exploit the deeper, lower grade ores. Similarly, extraction of metal from the sulphide ores presented a major problem. With the Meiji Restoration in 1868 and the rapid opening of the country to western ideas came a rejuvenation of the mining industry. Mining, milling, and smelting processes were greatly improved, resulting in the extension of development work into hitherto inaccessible portions of orebodies. The demand for metal during World War I stimulated activity throughout the industry; copper mining, particularly, expanded to a point where Japan ranked second among world producers.* During the depression following the war, mining underwent a severe contraction in activity, recovering slowly during the late 1920's only to undergo a further setback in the worldwide depression which started in 1929. Mineral output declined in 1930, but the beginning of Japanese expansion, 1931 to 1932, brought about an increase in the output of mine products starting between 1932 to 1934. Production continued to rise, the peak output being attained 1939 to 1943 in virtually all commodities. The greatest production of iron ore, chromite, and manganese was not experienced until 1944 when blocked imports, coupled with strong demands for steel, provided the needed incentive. Through necessity, Japan produced much of her mineral needs from her own mines at high cost during World War 11. Taking the 1932 to 1936 period as an average, however, Japanese mines produced only 16 pct of the iron available to industry, 8 pct of the lead, 33 pct of the zinc, and 68 pct of the copper. Nearly all of Japan's output of gold, silver, copper, lead, and zinc between 1925 and 1945 came from 30 to 49 pct of the mines for which production data are available; for example, 66 of 174 gold-silver mines accounted for 98 pct of the output of this group. Further analysis of mine output indicates that a large part of Japan's metal output was derived from a few large mines; about 10 gold-silver, 15 copper, 3 lead-zinc, 5 manganese, and 2 iron mines account for most of the production. Many of these mines have been operated for 200 to 400 years. With the

Jan 1, 1952

By S. Weinig, E. S. Machlin

IN a previous study of the grain boundary stress relaxation phenomenon,' the authors had arrived at the conclusion that two successive steps were involved in the complete relaxation of stress at a grain boundary. These two steps were grain boundary migration and grain boundary slip. It was further suggested that the slower of these two processes is rate controlling. In an attempt to ascertain the validity of the above, a study of grain growth kinetics on the same alloys used for the previous study was undertaken. Separate wire specimens, 0.030 in. diam, in the same as-deformed state (99 pct reduction in area) as in the previous investigation were heated to the various temperatures for a sequence of times. The compositions used and the temperatures of testing are summarized in Table I. A statistical determination of grain size was then made on each specimen, using the average length of traverse between boundaries as a measure of the grain size. Examples of typical results obtained are shown in Figs. '1 and 2. The grain growth law D = ktn was found to fit most of the data. Using this relation, values of the grain growth exponent n were calculated and the inter -esting dependence of n on solute content shown in Figs. 3 and 4 was then obtained. Also, it was found that the constant K which was obtained from extrapolation is roughly independent of solute content. (K is uncertain to within ±25 pct.) An estimate of the activation energy for grain growth was made using the technique of Burke? This method utilizes the plot of the reciprocal of time necessary for the grain size to increase from a definite starting size Do, to a final size D, for each isothermal heat treatment. The slope of this curve corresponds to the heat of activation for the process. In Fig. 5 the rate of grain growth is plotted against the reciprocal of absolute temperature. The activation energies were obtained from these curves and are summarized in Fig. 6. Discussion It is interesting to note that the same saturation in the variability with concentration has been found for the grain growth exponent n in this research as was found in the previous researches on both grain boundary stress relaxationa and internal friction of dislocation origin." The significant feature is that saturation in the variation of these properties has been found to occur at about the same level of solute concentration, namely, between 0.1 and 0.5 atomic pct. The only characteristic common to these different quantities is that solute adsorption at grain boundaries and dislocation can take place and affect these values in a similar fashion. Thus, it is concluded that the grain growth exponent n depends upon solute adsorption at the grain boundaries. It appears, therefore, that theoretical considerations leading to the contrary conclusions5,6 have not treated the real process by which solute atoms exert their effect on grain growth. Prior to the analysis of the observed activation energies, certain of the aspects of the aforementioned grain boundary stress relaxation investigation will be briefly reviewed. It was found that with the addition of solute atoms the grain boundary stress relaxation peak (internal friction vs temperature) decreased in intensity. Simultaneously a second peak was observed which increased with increasing solute atoms. Both of these peaks were shown to be the result of grain boundary stress relaxation. They have been termed, respectively, poire peak and solute peak. It was suggested that the solute peak was dependent upon grain boundary migration, whereas the pure peak was a manifestation of grain boundry slip. Activation energies for both peaks were obtained as a function of Solute content. The value of the activation energy in the temperature range corresponding to the solute peak was

Jan 1, 1958

By Pablo Hadzeriga

Great Salt Lake brine was subjected, in laboratory scale, to conditions simulating solar evaporation. Solid phases and the variation in composition of the liquid phase throughout the potassium salts crystallization zone are presented. These experimental results are discussed and compare with stable equilibrium data of the quinary system Cl-, SO4, Na+, K+, Mg + + and H2O. The Great Salt Lake brine in Utah is a vast reserve of several potential chemicals which has attracted much attention in the recent past. From the standpoint of an ore reserve, it may be said that the lake contains a low grade ore which has to be concentrated by evaporation of water. Because of the fairly favorable weather conditions in the Great Salt Lake area, it seems reasonable to consider solar evaporation as an economical means of concentration. From the view of phase chemistry, the main constituents of Great Salt Lake brine make it belong to the five component system: C1-, SO4=, Na+, K+, Mg++ and H2O. Because of their low concentration, the presence of other elements may be regarded as having little effect on the equilibrium of the five main components. Equilibrium data for the quinary system of our interest has been available for many years. Using these data, it is possible to calculate how the brine of the lake will concentrate when subjected to solar evaporation. In practicing solar evaporation, the weather factor and supersaturation conditions cannot be ignored, but their introduction into calculations using equilibrium diagrams is almost an impossible task. With the experimental work presented in this paper, an attempt is made to compare the stable equilibrium data of the quinary system as applied to the case of the Great Salt Lake brine and the dynamic equilibrium when that brine is subjected to continuous solar evaporation. Sea water brine belongs to this quinary system when only the major constituents are considered. The study of this system has been carried on since the last decades of the nineteenth century. The reason for these intensive studies was to give a geological explanation to the underground potash deposits. AS a result, many theories were presented trying to relate the equilibrium data with the geological formation, paragenesis and metamorphism of these deposits. Two main approaches were followed for the equilibrium studies of this quinary system. One of them, and the best known to us, was the school started in Germany in the last century by J. H. Van't Hoff. The other one, which started about the same time, was the school of N. S. Kurnakov in Russia. Both approaches were different and as a consequence generated opposite theories about the geological formation of potash deposits. The Van't Hoff school studied the quinary system of sea water under careful laboratory control, resulting in stable phase diagrams. During these studies, Van't Hoff recognized the difficulties in crystallizing some double salts.' ". . . . we unexpectedly found that some compounds whose formation in the solutions under investigation was possible at 25°C, nevertheless totally failed to put in an appearance. These, not to mention the calcium salts, were leonite, MgSO4 . K2SO4 4 H2O; kainite, KC1-MgSO4 . 3 H2O; and kieserite, MgSO4 .H2O. Even exceedingly slow crystallization with the addition of the compounds themselves did not remove the condition of supersaturation in the case of these bodies." However, those salts (leonite, kainite, and kieserite) were included in the stable phase diagram at 25°C as shown in Fig. 1 using D'Ans data.3 The figure also shows the sequence of crystallization of the Great Salt Lake brine according to the stable diagram. In later studies by the Van't Hoff school, especially by D'Ans,3 Karsten,4 Autenrieth,5,6 attention

Jan 1, 1968

By R. L. Hadley, G. Derge

The amounts of oxygen in liquid iron-titanium alloys up to 50 pct Ti were measured and the oxide phases in equilibrium with these alloys were determined by using TiO² crucibles. A minimum of about 0.003 pct 0 occurs at less than 1 pct Ti. The various liquid and solid oxide phases can be arranged on a ternary isotherm. THREE characteristics of titanium contribute to the interesting aspects of its reaction with oxygen in liquid iron: 1—It forms a number of stable oxides. 2—All oxides have large negative free energies of formation. 3—The pure metal exhibits high solubility for oxygen. In addition, the use of titanium as a stabilizing and strengthening element in alloys suggests a need for investigation of its reactions in amounts greater than those normally required for deoxidation. These higher compositions are also of interest because they occur during the addition of ferroalloys and produce deoxidation products before dilution is achieved. Titanium Deoxidation in the Literature While the use of titanium as an alloying element and deoxidizer has been common for several decades, the first detailed study of its deoxidizing characteristics was carried out by Wentrup and Hieber' in 1939. These authors determined the extent of deoxidation possible with titanium and identified, by microscopic examination, the products formed. They found that titanium in excess of 0.2 pct did not reduce the oxygen content of iron below 0.004 pct at 1600°C. They identified the spinel FeO-TiO² at low titanium concentrations and Ti2O³ in melts containing more than 0.2 pct Ti. A portion of the area surveyed by Wentrup and Hieber has been investigated recently by Evans and Sloman² using X-ray diffraction for inclusion identification. They concurred in general with the findings of Wentrup and Hieber, but reported the existence of Ti³O5 as a deoxidation product in the range near 0.1 pct Ti. Comstock, Urban, and Cohen³ discussed the deoxidation of iron with titanium, correlating the data of Wentrup and Hieber with free-energy data for the titanium oxides. Scatter at high titanium values was considered attributable either to the formation of an oxide exhibiting high solubility in liquid iron or to the entrainment of solid oxide particles leading to high apparent solubility. Chipman' has calculated the deoxidation curves for titanium based on free-energy data for titanium oxides. These calculations (Fig. 1) and those of Comstock, Urban, and Cohen involve assumptions as to the activity coefficients of oxygen and titanium in liquid iron, since such data have not been determined. Investigation of the deoxidizing characteristics of chromium" and vanadium0 have shown that these elements decrease the activity coefficient of oxygen in liquid iron, leading to a tendency toward relatively high oxygen values and experimental scatter in highly alloyed melts. Experimental This investigation involved two major objectives: 1—determination of the composition of the oxide phases in equilibrium with liquid Fe-Ti alloys, and 2—establishing the pattern of oxygen solubility in liquid Fe-Ti alloys at titanium concentrations in excess of that required for attainment of minimum oxygen content. Raw Materials Ingot iron was used having the following composition: 0.02 pct C, 0.017 pct Mn, 0.01 pct Si, 0.02 pct S, 0.006 pct P, and 0.03 pct 0. The titanium used

Jan 1, 1956

By P. Hudock

Thin films of PbS and CdS have been deposited on insulutiug- sapphire substrates by euzploying a subliwzation technique rising near-equilibriurum conditions in ultrahigh vacuum. Oriented polycrystalline PbS and CdS films on sapphire have mobilities approachitzg values characteristic of bulk crystals. In the case of CdS, special purification procedures were crucial in obtaining high-mobility films. By performing the depositiotz of a compound within a small chamber heated to a UNiform temperature within a Vacuum systerrz, it is possible to approach equiilibrium cod-tions r.~litl~irs the chamber. In this way; the sublimation and condensation proceed under near equilibrium co?zditions, and it was found that the properties of dePosited corrzpounds were similar to those of the source material. Oriented PbS films on sapphire have a roor?l- temperature unobility of 700 sq cm per v sec and a T-5/2 temperature dependence characteristic of bulk crrystnls. CdS layers on sapphive Izazse a resistivity ~uhich is corrlpamble to the resistivity of the souvce crystals over a range of four ordem of rrlagnitude. Irz contrast, Ihe resistivity of CdS films prepared by conventional evaporation depends largely upon the substrate terrzperat~ire and postheating treatrrletlt, vather than upon the properties of' the solirce rrraterial. In the case oj the CdS films the achievment of high carrier mobility required the cargful exclusion of cont~~~~ii~unts from the films. The steps taken to accomplish this included presublimation of the source material, performance of the deposition at a backg-m~ozd pressure less than 10-8 Torr, and a solvent extraction with liquid cadminum, similar to that described by Aven and woodbury. The deposition of ,films under near-equilibrium conditions is being op-plied to other classes oj COmpounds. II should he of 1:allce ,tor llle deposition of nny matevial that decomposes in the vapor state. THIN films of PbS and CdS have been deposited on insulating sapphire substrates by employing a sublimation technique using near-equilibrium conditions in ultrahigh vacuum. Oriented polycrystalline PbS and CdS films on sapphire have Hall mobilities approaching values characteristic of bulk crystals. In the case of CdS, special purification procedures were crucial in obtaining high-mobility films. Depositions were performed in a small quartz sublimation cell shown in Fig. 1 that contained the compound and the single-crystal sapphire substrate. The substrates were etched in tapered solutions of hot phosphoric acid, rinsed in deionized water, dried, and transferred to the cell under dry nitrogen. Graphite crucibles and substrates were used in experiments where it was necessary to provide an inert test vehicle for the analytical determination of impurities in the source crystals and deposited films. The crucibles and substrates were machined out of spectrographic-grade graphite having less than 1 ppm of impurities. A quartz mask was sometimes used to define a pattern which could be used for Hall and resistivity measurements. Close spacing between the compound and substrate was maintained to insure better transport of each component from the source material.

Jan 1, 1968

A Page Adam. H. F. and Ruggles! G. H., Results with Xanthate at Inspiration............. 337 Adnic proertiea of....................... 474 Africa; Gold Coast, manganese...........abs.474 191 mining districts......................... 254 Aid, H., biography........................ 154 Air compressors. In flotation................ 45 Air-gas lift: gae-oil ratios relation of......abs. 475 Gulf Coast. mechanical installations... .abs. 475 Seminole field........................abs. 476 Air-sand process. cleaning coal............abs. 89 Alabema, Birmingham district. drainage. . abs. 86 " Alclad." aluminum alloy................. 395 Alexander. W. Homer. biomaphy .......... 37 Allen. Andrews. Preparing Coal after Mechanical Loaders......:....T................. 181 Allison. F. H., Comparison of the Effect of Nickel and Cobalt in Steel.............aha. 151 Alloys: See Metals named. Aluminum: heat treatment...............abs. 444 machining...........................abs. 443 Norway............................... *465 Aluminum alloys: " Alclad "................ 395 application of........................abs. 443 deformation..........................abs. 150 Aluminum Bronze Manufacturers Institute. organized.............................. 188 "Alumnum. The Metal and Its Alloys." B.R. 42 Aluminum-iron-silicon alloys, equilibrium relations.............................aba. 444 Aluminum-manganese alloys, equilibrium relations in..........................aba. 87 Aluminum-silicon alloys. heat treatment. .abs. 398 American Association of Petroleum Geologists. annual meeting................... 81 American engineers. war memorial.......... 499 American Institute of Mining and Metallurgical Engineers: Committees: Douglas. James, medal...... 324 Geophysical Prospecting............. E. 204 Mining Medal........................ 324 Nominating: Election................. 235 report............................. 479 Constitution. votes for.................E. 494 Constitution and By-laws. draft of....... 374 Directors: Minutes of meeting.....110. 235. 323 publications.......................... 523 Directors adopt badge................... 324 Dues. increasing........................ 111 Dues are Due.......................... 236 Institute of Metals Division: annual lecture............................ 187 data sheets.......................... 187 Detroit meeting, (September). . .307. 353, 440 lecturer C. H. Desch................. 15 Local Section Committee. members..... 187 New York meeting (February)......... 115 Research Committee.................. 187 research problems..................... 6 Local Sections: Arizona............................. 365 Boston, meetings................40, 197. 486 Page American Institute of Mining and Metallurgical Engineers—Continued Chicago, meetings...40, 94, 156, 196,240,417 Chicago and St. Louis, joint meeting, Danville....................... 486, 527 Colorado, annual meeting.............. 285 Columbia, meetings 39, 93, 156, 196, 238, 286, 327, 365, 486, 527 El Paso, organized.................... 238 Fort Worth, organization meeting...... 485 Joplin-Miami Zinc, meeting.......... 286 Lehigh Valley: annual meeting......... 197 meetings...................... 239, 527 Los Angeles, meeting.................. 285 Mid-Continent, meeting............... 156 Minnesota, annual meeting............ 326 Montana, meeting.................... 40 New York, meetings................240, 284 North Pacific: meetings.... 39, 157, 197, 527 Fieldner, A. C, speaker............. 486 resolution.......................... 439 Oregon, officers....................... 196 Pittsburgh: meetings........... 156, 327, 526 mine transportation conference....... 485 St. Louis and Chicago, joint meeting, Danville........................... 486 San Francisco, meetings............... 92, 197, 239, 285, 439, 486, 527 Southern California, meetings 40, 157,197,283 Upper Peninsula, meeting.............. 486 Utah, meetings..........40, 93, 156, 239, 365 Washington, D. C, meetings 40, 93, 365, 528 Medals: Douglas, James................34, *84 Fritz, John, award, 1928.............. 480 Johnson, J. E., awarded T. L. Joseph... 83 Saunders, William Lawrence........... 34 meetings: Cleveland, April, 1927.......166, 207 Detroit, Sept., 1927............307, 353, 440 Fort Worth, Oct., 1927.........307, 353, 471 • future............................. E. 102 Knoxville, Nov., 1927.............. 480, 500 New York, Feb., 1927, announcements, programs, etc.................33, 50, *106 Salt Lake City, Aug...............306, *422 Members and jobs...................... 477 Membership, benefits.................8, E. 2 Officers and directors.................... *122 Open-hearth Conference................. 282 Petroleum Division: activities............ 186 annual meeting...................... . Fort Worth meeting............307. 353 471 luncheon............................ 91 New York Petroleum Club, officers..... 110 nominations........................23. 519 officers.............................. 117 President's trip. May.................281. 398 Publications: Directors discuss............ 523 Institute of Metals Division Volume... . 178 Petroleum Volume.................... 192 Technical Publications...........E. 294 323 Transactions. Vol. LXXIV. LXXV....00. 524 Review of the year.........:........... ...E. 104 Stockholders. report to..............E. 49

Jan 1, 1927

By Robert M. Lewis

Can wet-ground mica be produced from mica schist ores, and if so, would its properties be comparable with products now on the market? This was the problem which prompted the development of a flowsheet for the beneficiation of mica schist. The fine-ground mica industry relies on weathered pegmatites and alaskite ores as sources of raw material. The mica concentrates obtained from these ores are relatively clean and require moderate beneficiation to meet color and bulk-density specifications. Mica obtained from mica schist is iron-stained and decrepitated and requires additional beneficiation to produce a wet-ground mica product. Despite these adversities, there are several advantages to be gained by mining and processing mica schist ores. The mica content of these ores is considerably higher than that of the pegmatites and alaskite ores now being mined, often containing in excess of 40% mica. This higher-grade mica feed should promote a lower mining and milling cost. The mica schist ores with their higher mica content would mean less tailings to handle; however, there would have to be a means for disposing of acid and sulfate wastes. A flowsheet was developed whereby mica could be recovered from mica schist ores and beneficiated in a way that would result in a high-quality end product.' The essential phase of the process involves using reducing reagents to increase the magnetic susceptibility of iron contaminants which then can be removed with a magnetic separator. General Treatment Procedure The mica is floated by either of two methods, an amine float in an acid circuit' or the U.S. Bureau of Mines (USBM) amine-fatty acid float in an alkaline circuit' The flotation mica is ground in a pebble mill to liberate contaminated particles and expose iron-stained surfaces. The concentrate is then conditioned with a reducing reagent and passed through a Frantz Ferrofilter for the removal of magnetic material. The nonmagnetic product is acid-leached to further reduce the iron contamination. The leached material is then ground to product size and checked for specifications. Discussion of Flowsheet and Its Development Flotation: The mica is concentrated by flotation using either of two methods: an amine float in an acid circuit, or the USBM amine-fatty acid float in an alkaline circuit.' The writer does not wish to elaborate on the relative merits of the two flotation procedures; however, after investigating ores from various localities, it was felt that the amine float in an acid circuit was more consistent when using a standard procedure. Iron Reduction and Magnetic Separation: This phase of the flowsheet may be considered the key to the process. It was developed when it became apparent that hot acid leaching had just about reached its maximum effective¬ness for color improvement. S. C. Sun and William Hirsh in "Hydrochloric Acid Leaching of Iron From Pennsylvania Aluminous Clays,"` reported iron extraction for -20-mesh minerals of 82.20% for hematite and only 33.46% for magnetite. The test procedure employs a 19% hydrochloric acid solution, 104°C temperature, and 10% solids. It was felt that magnetic removal of some of the iron contaminants before leaching therefore would be advantageous. Since the iron minerals present as contaminants vary in magnetic susceptibility, it was decided to experiment with reducing reagents as a means of influencing their magnetic properties. Zinc hydrosulfite or sodium hydrosulfite were the reagents used throughout the tests. The procedure developed consists of conditioning the mica concentrate in a closed container at approximately 5% solids with a reducing reagent. Approximately 10.0 lb of zinc hydrosulfite per ton of mica concentrate is used. The ferric iron is reduced to the ferrous state (magnetic) and oxidation is retarded by isolation from the atmosphere in a closed container. After the desired conditioning time has elapsed, the material is passed through a Frantz Ferrofilter where the magnetic particles are retained and discarded as

Jan 1, 1972

By Robert B. Hall, Charles W. Bauer

Alunite, KA13(SO4)2(OH)6, was used from the 15th century until early in the present century as a source of potash alum and aluminum sulfate. Since early in this century its use for this purpose has declined to virtually nil with acid treatment of bauxite or clay replacing alunite as a source of alum. During the First World War alunite served as an emergency source of potassium sulfate fertilizer in the United States and Australia. More recently, alunite has been investigated as a nonbauxite ore of aluminum, with potassium sulfate and sulfuric acid as recoverable byproducts. The Soviet Union established a commercial-scale plant in the mid-1960s producing cell-grade alumina from alunitic ore and recovering byproduct potassium sulfate and sulfuric acid. Interest in nonbauxite ores, including alunite, is expected to continue in countries that presently are dependent on imports of bauxite or alumina to support their aluminum industries. The US has the largest aluminum producing and fabricating industry in the world but imports more than 90% of its aluminum-bearing raw materials. This chapter reviews the status of alunite as a potential source of aluminum with byproduct fertilizer, and the resources available to support this industry, whenever this becomes economically feasible. Historical Background Alunite has received little attention until recently, yet for centuries it held a higher place than many of the industrial minerals considered essential today. The deposits of alunite at Tolfa, northwest of Rome, were mined almost continuously from 1462 until about the middle of the present century (DeLaunay, 1907; Lombardi, 1977). The potash alum made from the alunite, called Roman alum, was a valued commodity in international trade, being used in the making of paper, in tanning leather, and as a mordant in textile dyeing. The industry at Tolfa is said to have been controlled by the Vatican and was a source of revenue for the Papal State until fairly recent times (Lombardi et al., 1977). Since early in the present century, the preferred method for making alum has been by acid treatment of bauxite, clay, or other aluminous material, so that alunite has fallen into near-total disuse as a source of alum chemicals. Appreciable quantities of mixed alunite-kaolin rock still are quarried in the Tolfa district for use by the Italian cement industry (Lombardi and Mattias, 1979), but the role of this material in the making of Italian cement is not clear. A large deposit of alunite-bearing rock in Chekiang province, mainland China, is said to have yielded a "great quantity" of alum over a period of several centuries (Yih, 1931). During the 19th century, alunite deposits in France, Spain, Australia, and other countries were exploited as raw material for making potash alum and aluminum sulfate. Interest in alunite was revived during the First World War, not for alum, but as an emergency source of potassium sulfate fertilizer. Large veins of nearly pure alunite had just been discovered on Alunite Ridge, 11 km southwest of Marysvale in Piute County, Utah (Butler and Gale, 1912), and these were quickly brought into production for making potash fertilizer to replace supplies cut off from the famous deposits at Stassfurt, Germany. An estimated 225 kt of Marysvale alunite was mined and processed at that time (Callaghan, 1973). Similarly, Australian alunite deposits were mined as a source of potash fertilizer during both world wars (Hall, 1978).

Jan 1, 1983

By Kenneth A. Gutschick, Robert S. Boynton

Lime has become a very general loosely used term that includes virtually all finely divided types of limestone as well as burned forms of lime. However, actually (and according to Webster) lime is only calcined limestone (known as quicklime, calcium oxide, or calcia). It also embraces the secondary product of quicklime-namely hydrated lime or slaked lime (calcium hydroxide). Fundamentally there are two basic types of limestone-high calcium and high magnesium (dolomitic). The dolomitic quicklimes and hydrated limes correspond to their "high calcium" counterparts in terminology. The only difference is that the dolomitic types are a combination of the elements, calcium and magnesium, whereas high calcium limes contain only 0.5 to 2 pct of magnesium. Both types, which are made through calcination at high temperatures, are shown chemically as follows: [ ] In the above reactions either type of limestone is burned (calcined) in furnaces known as lime kilns so that the carbon dioxide content of the stone is expelled as a gas. With high calcium limestone 44 pct of the weight of the stone is lost; with dolomitic limestone the loss is 48 pct. Also, both reactions are chemically reversible since quicklime absorbs carbon dioxide readily and in so doing transforms itself back to the original carbonate form. Most quicklimes correspond in size to the original rock, which varies from one to eight inches, depending upon the type of kiln employed. Slight shrinkage occurs during calcination. A few limestones when calcined will decrepitate (break down) into a fine powdery form; however, chemically it is the same as the pebble or lump-sized material. For some uses the lumps of quicklime are pulverized and packaged. Since quicklime has great affinity for moisture and carbon dioxide, it is perishable. When it absorbs moisture and CO2 slowly from the atmosphere, the resulting product is "airslaked." Consequently, it must be stored in dry, moisture-proof areas that are free from CO2. Because of its fugitive nature, it is usually consumed within a few weeks after manufacture. A more stable form of lime is hydrated lime-commercially made by adding water to the quicklime until the lumps break down into a fine, white powder. This product, which is usually packaged, can be stored more safely; and many of its uses are interchangeable with quicklime. The irreversible chemical reactions of hydrated lime for both high calcium and dolomitic types are: [ ]

Jan 1, 1960

By Thomas C. Patton, Eric S. Cheney

L-D gold mine is 3 miles south of Wenatchee, central Washington. Recognition of locally mappable conglomerates, sandstones, and shales within the Paleocene (?) Swauk formation led to the discovery that the known epithermal gold mineraliza¬tion is limited to a previously unknown imbricate thrust zone of unknown displace¬ment cutting the open folds o f the area. Hypabyssal intrusions also are most nu¬merous along the trace of the thrust faults. Yakima basalt of Miocene age uncon¬formably overlies the Swauk formation, thrust faults, and the intrusions. The L-D mine is in a northwesterly trending imbricate fault block of nearly vertical strata. Microscopic native gold, electrum, and naumannite occur in quartz veins and as low¬grade disseminations in the well-silicified portions of arkosic sandstones. Fault gouge on the "footwall fissure" limits ore to the hanging-wall block of the fault. Epithermal mineralization and alteration are spatially related to the nearby intrusions. Right lateral north-south faults offset the ore a maximum of 300-400 ft along any one fault. From 1949 until January 1967, when the L-D mine closed, 1,036,572 dry tons of ore valued at $14,962,305 and averaging 0.396 oz Au per ton and 0.607 oz Ag per ton were mined. Recognition of the structural control of the mineralization and altera¬tion within the Swauk formation suggests several areas within and outside of the mine that warrant further exploration. For example, the high-angle faults, imbricate thrusts, and the intrusions of this area are on strike with similar features along the major Entiat fault zone 10 miles to the northwest. The L-D gold mine (previously called Squilchuck, Wenatchee, Golden King, Gold King, and Lovitt) is located in Squilchuck Canyon about three miles south of Wenatchee, Chelan County, Wash. (Fig. 1) on the eastern edge of the Cascade Mountains. The major stratigraphic unit in the area is the nonmarine early Tertiary Swauk formation, which is predominantly com¬posed of arkosic sandstone. The lack of laterally per¬sistent units and the paucity of outcrops heretofore have precluded all but the grossest subdivision of the Swauk.' As a result, the true significance of the open folds in the Wenatchee area was not recognized, and although the geology within the mine was known, the structural setting of the mine was virtually unknown. During the present investigation, several locally map¬pable units were discovered within the Swauk which, in conjunction with mine maps, indicate an imbricate thrust zone. The epithermal gold mineralization, most of the hypabyssal intrusions, and the zones of hydrother¬mal alteration are located within or near this thrust zone. The purpose of this paper is to: (1) to describe this structural setting, rather than the detailed mine geology already presented by Lovitt and Skerl,2 and (2) to sug¬gest that from an exploration standpoint, this struc¬tural setting may be repeated in areas inside and outside of the mine area. Stratigraphy and Structure of the Swauk Formation General: The Swauk formation contains three distinct lithologies. Typical Swauk sandstone is gray on fresh surfaces, poorly sorted, weathers tan, and is cross¬bedded; it consists of 40-60% quartz, 20-40% feldspar, and up to 10% muscovite and biotite 9 The typical shale is gray to dark brown, very soft, and interbedded with silty sandstones; sharp contacts exist between the

Jan 1, 1972

By C. E. Gregor

Current modern trends in mining show that hoisting shafts are being expected to fulfill an important ventilation function. However, where rigid guides and supporting structures are mandatory, ventilating power costs will be high, unless the transverse structures (shaft sets) are streamlined. In the complex configurations of hoisting shaft furniture patterns, either for original design purposes, or for the adjustment of an existing shaft, a useful degree of aerodynamic optimization, by reduction of overall resistance, can only be effectively evaluated by model studies in a special purpose wind tunnel, as described herein. The author has shown that, in one particular case, an annual power cost saving of more than $50,000 was indicated, mainly by the selection of buntons and dividers of improved aerodynamic profile. Mining companies are now realizing the advantages that can accrue through proper attention to the aerodynamic aspects of mine shaft design. With progressive demands of civilization for an ever greater supply of coal and metals, and with greater depths of working, higher rates of extraction, advanced degrees of mechanization, and improved standards of environmental health and work comfort, shaft design is becoming an extremely important aspect of mining operations. In addition, because the cost of sinking, equipping, maintaining, and operating a mine shaft is continually increasing, there is a greater incentive to use a shaft for the maximum duty, consistent with economy. The current natural tendency is therefore to sink multi-purpose shafts, to install as much equipment as is safely and effectively possible to operate; and then, over and above all these functions, to use the shaft to its maximum possible ventilating capacity, as a major, or auxiliary, airway. Clearly, this latter additional function can only be employed at a significant cost, since the most economic operating airways are smooth-lined and devoid of internal rigid furniture. It is true that hoisting functions can be carried out in such shafts if rope guides are used, but only under restricted conditions. Where a manway must be provided, fixed guides and supporting structures are mandatory, and ventilation costs become accentuated. It is generally recognized that the prime element of cost in ventilating a mine can be represented by the power required to drive the fan unit. It is also known that the largest component of total mine resistance is that of the downcast shaft, where this shaft is equipped with rigid guides for hoisting purposes. The pressure loss varies directly as the resistance of the shaft, in accordance with a modification of the Atkinson equation p = RQ2/3600. Furthermore, the air horsepower drawn is given by the equation AHP = (p x 0/33000. Therefore, the power required to drive the fan unit varies as the resistance of the shaft and the cube of the air quantity. Clearly, if Q is a specified minimum, then the only way to minimize the pressure loss (and the power costs) is to reduce the shaft resistance R. (See Table I for a list of symbols.) The resistance to airflow of a shaft equipped with rigid guides and supporting structural sets consists of two components which, for the sake of analytical simplicity, can be regarded as additive. 1) Frictional resistance, represented by the viscous frictional forces within the boundary layer, as air flows along the surfaces of the walls and the periphery of the vertical elements (guides, pipes, cables, ropes, etc.) of the furniture. The friction loss can be readily calculated by the Atkinson equation, after selection of a suitable value for the friction factor K from the McElroy compilation.' KxLxO xQ2 5.2 xA3 Friction loss can be minimized by lining the shaft walls (particularly if of circular or elliptical cross-section) with a smooth continuous lining of monolithic concrete.

Jan 1, 1968

By H. E. Wilborg, G. V. Henderson

This chapter deals with those sands employed by foundries for the manufacture of cores and molds used in the casting of such common metals as steel, gray iron, ductile iron, aluminum-based alloys, and copper-based alloys. Most metal is cast in "green sand" molds in which the sand mixture, consisting primarily of silica, clay, and water, forms a plastic material. This plastic material is rammed around a pattern in a split box called a flask, then the box is separated, the pattern is withdrawn, and the metal is poured. These green sand molds are broken apart when the metal has solidified. The sand is then remixed and a new mold made for the next casting. To a lesser extent silica sand is mixed with resins or oils and, after forming, the mixture is hardened by either heat or chemical reaction. In any case, the foundry mold or core is used just once before being broken up and remixed for manufacturing another core or mold. By far the greatest consumption in the foundry is of those types of sand high in silica, low in clay (under 0.5%), and occurring in numerous locations thoughout the United States. Consumption End Uses Foundry sands fall into four distinct categories: silica sand, lake sand, bank sand, and natural molding sand. Silica sand is a general term used to describe washed, graded, dried, and cooled clay-free sands which are found in many parts of the United States. They generally contain 90 to 100% silica but vary greatly in chemical analysis and grain structure. Lake sand is the sand dredged from Lake Michigan or dug from the dunes bordering that lake. Bank sand is composed primarily of silica grain but containing up to 5% clay. It is shipped in both the damp and dry state but the normal processing of bank sand consists only of drying and scalping. Occasionally two different grades of bank sand are blended prior to or during loading. Natural molding sand is a term used to describe a sand containing sufficient clay as mined to be suitable for molding purposes. Clay content may vary from 10 to 24% with the balance of the material consisting primarily of silica sand. These sands range in fineness from one peaked on the 20 mesh to one which contains more than 50% of the particles passing the 270 mesh. Lake sands are relatively coarse and are used in great quantities for the production of automotive castings. Bank sands and natural molding sands are gradually phasing out as the production of castings becomes more a science and less an art. The vast majority of foundry sands used today are those known as the silica sands and lake sands. Foundry sands are used in two basic ways. The majority are used in the green or uncured state with the balance used in the cured or set state. Green sand is grain bonded with clay and contains varying amounts of water to plasticize the clay in order to bond the grains together. Green sand is either naturally bonded or "synthetic" sand. The naturally bonded molding sands are generally found in the Hudson River Valley of New York State, Cumberland County in New Jersey, and in Tennessee and Mississippi. Synthetic sand is molding sand compounded in the individual foundry. It is made up primarily of graded silica sand, western or southern bentonite, water, and carbonaceous and/or cellulosic materials. Green sand is used primarily in making molds. The use of naturally bonded green sand has been declining for many years, primarily because naturally bonded sands are more difficult for the foundryman to find and control. Synthetic green sand, on the other hand, can be

Jan 1, 1983

By W. G. Pfann

Under certain conditions, a molten zone can be made to move through a solid by impressing a stationary temperature gradient across the solid. This phenomenon can be utilized in fabricating semiconductive devices, growing single crystals, joining, boring fine holes in solids, measuring diffusivities in liquids, small scale alloying, and purification. Fundamentals and exemplary applications are outlined. ANEW aspect of the travel of a molten zone through a solid is considered in this paper. Whereas, in previously described zone melting techniques,'-' zones were usually caused to move by changing the position or temperature of a heat source, in the present technique a zone is made to move by impressing a temperature gradient across it. For example, a thin layer of molten A1-Si alloy, sandwiched between two silicon slabs having a temperature gradient normal to the layer, will travel through the hotter slab. One feature of temperature gradient zone melting, as this technique will be called here, is that zones of unusually small dimensions can be maintained. This leads to such diverse uses as have been listed in the abstract. In this paper fundamentals of the movement of a molten zone in a temperature gradient, and some practical applications thereof, are outlined. Assumptions The movement of a molten zone through a solid body of solvent substance in which a temperature gradient exists will be discussed. The length of the zone will be its dimension in the direction of motion. The direction of the temperature gradient will be toward the region of higher temperature. To be considered is a binary or higher order solute-solvent system in which the concentration of solute in the molten zone is sufficient to produce a lowering of the freezing temperature of the solvent which is at least of the order of the temperature range impressed on the charge. While we discuss primarily molten zones in a solid matrix, the principles are applicable to solid or vaporous zones in a solid matrix, a requirement being that the diffusivity in the zone be substantially greater than in the matrix. Diffusion of solute into the surrounding solid will be assumed negligible. In general, its presence will not basically alter the conclusions to be drawn. For illustration, a system represented by the partial constitutional diagram of Fig. la will be discussed, in which k, the distribution coefficient defined in the figure, is constant and less than unity. However, the technique is applicable to any solute- solvent system in which one component lowers the melting point of another component. Movement of a Molten Zone in a Temperature Gradient Consider, in the system AB of Fig. 1, where B is designated the solute and A the solvent, the effect of sandwiching a thin layer of solid B between blocks of solid A and placing the whole in a temperature gradient such that the temperature of the layer is above the lowest melting temperature of the system. Surrounded by an excess of A, the layer will dissolve A, becoming molten, and expand in length." As so- lution of A continues, at both interfaces, the mean solute concentration in the zone will move to the right in Fig. la, until at temperature T, the liquid at the cooler interface reaches the liquidus concentration C,. Solution of A thereupon ceases because the liquid T1 is saturated with A. The liquid at the hotter end of the zone, being at temperature T2, is not saturated with A at concentration C1. Hence solution of A continues there and concentration C2 is approached. A concentration gradient therefore is set up in the zone, causing A to diffuse toward the cool end and B toward the hot end. As a result, the liquid at the cool interface becomes supersaturated and a layer of crystalline A containing concentration kc, of B in solid solution freezes. Since a source of A

Jan 1, 1956

By Paul J. Sirkegian

A detail of the guide rollers is shown on Fig. I; and Fig. 2 shows a photograph of a guide roller that has been in service for several months on the haul-back cable. The former practice of supporting and guiding the haul-back cable on slusher hoists at Kimberly was accomplished by using the conventional type of snatch-block pulley, but it was found that an excessive amount of cable whipping and unreasonable wear on both rope and pulley resulted from this type of cable support. The guide, which is wholly made in the Corporation's shops, consists of two vertical and one horizontal roll assembled in a frame that is lagged to the cap timber in a slusher drift. In some slusher drifts more than one guide-roller assembly is used, depending upon the length of the drift. The installations usually are placed at 30-ft. intervals. We have found by actual experience that these rollers have increased the life of the haul-back cable approximately 200 per cent, as compared with the usage of the conventional sheave-block pulley. Cable Care The underground rigging crew and the slusher operators are trained in rope splicing. No cable clamps or knots are permitted on slusher cables and all cable "eyes" are spliced. The standard practice has been to use preformed cables of ?-in. diameter for haulage and ½-in. diameter for the haul back. The preformed cables have proved to be important factors for safety in use and speed in making splices. No external lubrication is used on the cables. It has been found that dry ropes have about 400 per cent longer life than those that were lubricated under these operating conditions. An accurate record is kept of all cable life and the tonnage moved by the slushers in each working place, and if the rope wear is excessive in any particular working place, an investigation is made to determine the reason. Wire-rope Sling for Snatch-block Anchor The rope sling is shown by Figs. 3 and 4. These slings support the snatch-block pulley for the haul-back rope. The standard practice at Kimberly is to use a roller-bearing snatch-block pulley 10 in. in diameter. The snatch-block housings are equipped with a clevis attachment in preference to swivel hooks, as it has been found that the more rigid clevis keeps the cable in better alignment. The snatch blocks are hung from the cable sling, which is wrapped around the posts of the drift sets as shown on Fig. 3. This method of anchoring the pulley provides better anchorage and flexibility than those of the rigid or spring-mounted types. When the slusher-drift headings are being advanced, the snatch blocks are hung from a "gin" pole ahead of the last set of timbers' An employee is assigned on each production shift to take care of the lubrication of all slusher hoists that are being used and he is provided with tools for minor repairs.

Jan 1, 1946

By Henry N. McCarl

Lightweight aggregates include a variety of mineral and rock materials used to provide bulk in concrete building units (block), light- weight structural concrete, and precast concrete units, as plaster aggregate, insulating fill, and other structural and/or insulation purposes. They are distinguished from other mineral aggregate materials by their lighter unit weight- 80 to 100 pounds per cubic foot (pcf) or less for loosely packed lightweight aggregates as compared to 125 pcf or more for crushed stone, sand, gravel, air-cooled slag, and other more common mineral aggregates. The various lightweight aggregate materials may be classified into four groups as follows: 1) Natural Lightweight Aggregate Materials-prepared by crushing and sizing natural rock materials such as pumice, scoria, tuff, breccia, and volcanic cinders. 2) Manufactured Structural Lightweight Aggregates-prepared by pyroprocessing clay, shale, or slate in rotary kilns or on traveling grate sintering machines. 3) Byproduct Lightweight Aggregates-prepared by crushing and sizing foamed and granulated slag, cinders, and coke breeze. With the exception of processed slag, this group is becoming less important as time passes. 4) Manufactured Insulating Ultralight- weight Aggregates-prepared by pyroprocessing ground vermiculite or perlite. This chapter will focus on the use of natural lightweight aggregates and manufactured insulating ultralightweight aggregates, and both the production and use of manufactured structural lightweight aggregates with some comments on byproduct materials. Detailed discussion of the production and use of both expanded and air-cooled slag is presented in a separate chapter in the utilization section of this volume, and there are separate commodity chapters on vermiculite, perlite, pumice, and volcanic cinders. Discussion of the commodities covered elsewhere in the book will be limited to observations on their utilization and relation to other lightweight aggregate materials. All materials considered as lightweight aggregates are derived from some initial natural raw material. While historical definitions have included such organic materials as straw and peat as lightweight aggregates, the discussion in this chapter will restrict itself to lightweight aggregates derived from mineral or rock materials. Thus while the actual lightweight material in use may be a manufactured product, it is derived from an ore or raw material that is obviously a mineral resource. In this context, then, the reserves of raw materials for lightweight aggregate use may include iron ore and flux stone (the components of blast furnace slag) as well as coal (the raw material from which cinders are derived) as well as those clay, shale, and slate deposits that display the characteristics of bloating or controlled expansion when subjected to elevated temperatures (1600 to 2400°F.).

Jan 1, 1975

By Orville R. Lyons, A. C. Richardson

DURING the past few years there has been a growing demand by coal operators for detailed information about the performance characteristics of the various dryers now being manufactured, preferably in such a form that the information could be used to determine the type of dryer best suited to a particular drying problem. A few operators have been desirous of going even further, because they are of the opinion that present-day dryers are not properly designed and that fundamental studies of dryer design should be undertaken. This growing demand for a method of dryer evaluation and the indicated need for fundamental research resulted in the sponsorship by Bituminous Coal Research, Inc. of a project at Battelle Memorial Institute to investigate the thermal drying of coal. In investigating this problem, the most feasible method of attack appeared to be a survey of active operations to determine the present status of the thermal drying of coal. Data were obtained, usually by actual plant tests with the cooperation of the various preparation engineers and their staffs, for Q screen-type dryers, 8 rotary- type dryers, 3 continuous-pallet-type dryers, 2 flash-type dryers, I tray-type dryer, and I cascade-type dryer. The testing program provided: (I) data for 24 coal- drying operations, (2) information about dried, (3) information about the previous treatment that the coal had received, and (4) information about operating costs. The data obtained through the dryer survey and relationships determined from the data make it evident that all of the types of dryers included in the survey can be operated with the same degree of effectiveness if each type of dryer is used to dry the sizes of coal for which it is best suited. It should be noted that it is usually undesirable and often impossible to dry different sizes of coal in one type of dryer. By use of the data in the report, any prospective dryer user, having determined the surface moisture percentage and screen- size analysis of the coal to be dried, the tons of water to be removed per hour, and the final surface moisture percentage desired, should be able to determine with a fair degree of accuracy: (I) the difficulty of the drying problem, (2) the approximate cost of drying, (3) the required dryer inlet gas temperature and volume and exhaust gas temperature and volume, (4) the type of dryer to use, and (5) the size of dryer required. The method of testing the dryers and determining the difficulty of drying can be used by operators already having dryers to check the performance of their particular dryer and determine whether they are operating effectively. Considerably more information was obtained from the survey of thermal drying than was anticipated when the survey was first considered. However, a great deal of work still remains to be done in the field

Jan 1, 1948

By K. H. Adams, T. Vreeland

The basal stress-strain behavior, dislocation density, and strain vale sensitivity of the flow stress uqere measured at room temperature on single crystals oj-zone-refined zinc , 99.999 pct Zn, and zinc containing 0.0025 and 0.02 wt pct Al. No significant change in the stvain rate sensitivity of the flow stress occurs with the addition of aluminum. Changes in dislocation substructure and nonbasal dislocation density were obserrled with the addition of aluminum. The effect of aluminum on the basal slress-strain behavior is explained in terms of changes in nonbasal dislocation density which determines the separation distance of junctions between basal and nonbasal dislocations. The onset of basal slip is associated with the breaking of attractive junctions. The chunge in basal dislocation density was found to be proportional to the one-third power of the plastic shear strain, indepedent of purity. THE effect of impurities in solid solution on the stress-strain behavior of single crystals has been the subject of many theoretical studies. These studies have generally concentrated on the interactions between impurities and dislocations to explain the hardening effeet of impurity additions. The theories that have been proposed include those of yott and Nabarro,&apos;l&apos; Cott-re11; Suzuki; and Fisher. In practice, it is generally difficult to predict the relative importance of the vari- ous effects due to impurities. In addition, secondary effects of impurity additions such as induced changes in dislocation density and substructure in the as-grown crystals have received little attention. Many experimental investigations have been conducted on the effect of impurities on the stress-strain behavior of fcc crystals. Hcp crystals have not received much attention except for the extensive work of Levine et a1.= on magnesium alloy crystals. Stuwe~&apos;&apos; has studied the effect of copper impurity on the critical shear stress of zinc crystals and has interpreted the results in terms of the impurity atmosphere model of Cottrell and Jaswon. This report presents the results of an experimental study of the effect of aluminum additions to zinc on the basal stress-strain curve and the strain rate sensitivity of the flow stress. Etch pit techniques were used to measure and characterize the dislocation density and substructure in the zinc alloy crystals. EXPERIMENTAL PROCEDURE Specimen Preparation. Compression test specimens of four different purities of zinc were prepared from CP grade zinc of 99.999 pct purity obtained from the New Jersey Zinc Co. An analysis of this material furnished by the supplier indicated the following impurities in wt ~ct: Pb = 0.0002; Fe = 0.0002; Cd = 0.00005; other <0.60011. A quantity of material of this grade was further purified by zone refining. Zone refining was carried out under a helium atmosphere by passing a double &apos;One furnace Over a 6-ft-10ng charge at the rate of 2 in. Per hr. * total of ten double zone passes were made before the charge was removed from the furnace. The first one third of each charge was used to grow single crystals. Quantities of two

Jan 1, 1969

By J. W. Gwinn

The underground haulage system in the lead-silver mine of the Bunker Hill & Sullivan Co., situated at Kellogg, Idaho, is the most extensive in the Cæur d&apos;Alene district, comprising about 35,000 ft. of tunnels, drifts, and cross-cuts where locomotive haulage is in use. The main working adit, known as the Kellogg tunnel, or No. 9 level, is approximately 10,000 ft. in length from the portal to the main working shaft and extends beyond that about 4,500 ft. in each of two directions. On this level all the ore from the Bunker Hill & Sullivan, the Sierra Nevada Consolidated, and the Caledonia Mining Co.—an average of 44,600 tons per month—is hauled to the mills, located a short distance from the portal. Electric haulage by the trolley-type locomotive is used on this level throughout. Below this level, at intervals of 200 ft. vertically, are five other working levels (Nos. 10, 11,12,13, and 14); and the total haulage distance on these, where electric locomotives are used, is approximately 16,000 ft. No. 13 level has but recently been opened up and is not working to capacity; while the main cross-cut on No. 14 has just been started, and has no need of power haulage. Of the four levels equipped with electric haulage, Nos. 11, 12, and 13 have storage-battery locomotives, and No. 10 still retains the trolley type by reason of the small and decreasing tonnage produced. Electrification - Electric power is supplied to the mine by two systems. The first, the 500-volt direct-current trolley, was installed in 1897 when the Kellogg tunnel was started, and was extended as the tunnel advanced, being used for the first trainload of ore on Nov. 17, 1902, since which time it has been in continuous service. The current for the line is transmitted on a No. 00 copper wire as far as the shaft, and on No. 0 wire throughout the remaining length of the Kellogg tunnel; and until the installation of the storage-battery locomotives it was used in practically all the cross-cuts and drifts on each level below No. 9.

Jan 1, 1915

By E. L. Dougherty, J. W. Sheldon

Using the numerical techniques shown in this paper it is possible to compute the simultaneous dynamical behavior of multiple fluid-fluid interfaces in two dimensions. Hence, fluid-fluid inter face models of several oil recovery processes can be constructed which allow prediction of the effect of fluid flow in two dimensions on the behavior of these processes while taking into account various physical effects, such as saturation gradients, phase changes and thermal stimulation. Techniques for constructing fluid-fluid interface approximations for several oil recovery processes are reuiewed. The validity of these techniques is established by comparison to experimental results. The availability of a computer program for computing the behavior of multiple fluid-fluid interface problems makes the use of potentiometric models uneconomic. The results indicate the areal sweep efficiencies obtained from potentiometric model studies at high mobility ratios may be somewhat optimistic. The results also indicate that the areal sweep efficiency curves obtained at high viscosity ratios 2—4 using miscible fluids in porous media are misleading and that considerable care must be taken in defining the mobility ratio for a displacement process, especially for miscible fluids. INTRODUCTION Many of the methods used to estimate areal sweep efficiency assume that the oil recovery process can be represented by one or more fluid-fluid interfaces. When the mobilities of the displaced and displacing phases are assumed equal, this quantity can be computed using analytical mathematical methods.l.2 For non-unit mobility ratio, experimental methods based upon the fluid-fluid interface concept include potentiometric and electrolytic blotter models. Scaled models of porous media have also been used to study this problem. Fay and Prats3 and Aronofsky4 used numerical techniques to compute breakthrough sweep efficiency, but neither of these studies were extended beyond breakthrough. Numerical techniques which have been developed for solving a general fluid-fluid interface problem are presented in a companion paper.5 These techniques have been incorporated into a program for a large-scale digital computer which can treat simultaneously the dynamical behavior of as many as six fluid-fluid interfaces. Thus it is possible to simulate the behavior of several oil recovery processes taking into account such effects as saturation gradients. Techniques for representing several oil recovery processes by fluid-fluid interfaces are presented. Processes considered are conventional water flooding with and without a mobile gas phase, hot water flooding, miscible flooding and enriched gas drive. Other processes, whose effects could be studied by this technique, but are not considered in detail here, are underground combustion and steam injection processes and various slug processes such as gas-water injection and alcohol-solvent processes. Theoretical results obtained with single interfaces for mobility ratios of 0.25, 1, 2, 4, 8, 16, 32 and 64 in a repeated five-spot are compared to experimental results for this geometrical configuration obtained using miscible fluids and conductive analogs. Predicted recovery curves for water flooding in a repeated five-spot are compared to experimental results reported in the literature. Example calculations are shown for a conventional water flood in the presence of a free gas phase and for a hot-water flood. MATHEMATICAL PRELIMINARIES We assume that the rock matrix is homogeneous with constant thickness, absolute permeability and porosity. The fluids are incompressible. We also neglect the effects of gravity.

Jan 1, 1965