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By Fathi Habashi

Chalcopyrite can be readily leached with HNO3, but the recovery of elemental sulfur does not exceed 50% under the most favorable conditions. Preheating at 500°C in a slightly reducing atmosphere increases the recovery of elemental sulfur to about 65%. Chalcopyrite can also be treated with HNO3 in a pressure reactor at 110°C to solubilize selectively the copper while iron remains in the residue as an oxide. The recovery of elemental sulfur also improved when preheated chalcopyrite is treated in this way. Stainless steel should be an adequate material of construction.

Jan 1, 1974

To assist the taxpayer of the oil and gas industry in correctly and expeditiously preparing his Federal tax returns, the Bureau of Internal Revenue has prepared a "Manual for the Oil and Gas Industry Under the Revenue Act of 1918." Although the endeavor has been made to anticipate all questions that may be asked, it is recognized that such a manual is only general and cannot cover all cases that may exist. The Manual is based largely upon information gathered during the fall of the year 1918 by a corps of geologists, technologists, and engineers. The investigation was undertaken primarily to furnish a basis for arriving at valuations, and depletion and depreciation deductions in connection with oil and gas properties. All fields in the United States were canvassed. Records of production of thousands of properties were collected and tabulated. These were carefully classified and studied by the most competent and experienced men in the country and the average future production curves and tables of valuation data were produced as a result of this study. Be American! Work and Save. Buy W. S. S.

Jan 7, 1919

SOUTHERN CALIFORNIA LOCAL SECTION Executive Committee THEODORE B. COMSTOCK, Chairman SEELEY W. MUDD, Vice-Chairman FREDERICK J. H. MERRILL, Secretary-Treasurer, 631 Higgins Bldg., Los Angeles, Cal. A. B. CARPENTER C. COLCOCK JONES A meeting of the Southern California Local Section, in honor of the presence of Secretary Bradley Stoughton, was held at luncheon in a private dining room of the Sierra Madre Club on Friday, May 21, 1915, 27 members and guests being present. In the absence, through illness, of the Chairman, Dr. Comstock, the vice-Chairman, Seeley W. Mudd, presided. After the luncheon, Secretary Stoughton made a short address on the activities of the Institute. George S. Rice, of the U. S. Bureau of Mines, who was present, then spoke on the activities of his Bureau. Local Secretary F. J. H. Merrill gave a description of the activities of the Southern California Section and its progress, and reported the latter as substantial and its condition as very good. After some informal remarks by other members present the meeting adjourned. F. J. H. MERRILL, Secretary.

Jan 7, 1915

By R. N. Thurston, E. A. Nesbitt, G. Y. Chin

A general relalionship between the amount of glide shear (due to slip) and the macroscopic shape change has been developed. Since the deformation can be large, finite strain analysis is employed. In this treatnzent, the shape change is expressed by a deformation gradient matrix, F - [axi/aXj1, where X and x refer to the initial and final positions of a particle. This matrix is readily evaluated in terms of the amounl of glide shear, a , and the direction cosines of the slip-plane normal and the slip direction for a single active slip system. In the case of deformation from slip on several systems, the product F(a) of the several deformation gradient matrices is first calculated. Then, by assuming that the final configuration is reached by a long series of small shears of magnitude a, occurring more or less alternately in the several slip sys- IN problems of crystal plasticity, it is often necessary to relate the amount of glide in the operating slip systems to the macroscopic strain components. This relationship is relatively simple when only small strains are considered. In this case the separate strain contributions from several slip systems are additive. Such a procedure is incorrect in the case of large plastic deformation. To the authors' knowledge, no general treatment of the latter problem has appeared in the literature. In specific instances, Mark, Polanyi, and schmid3 have derived the relationship between glide and axial elongation during tensile pulling of a single crystal when only one slip system is active. A similar relationship for duplex slip was worked out by v. Goler and Sachs.4 Taylor and Elam5 have also studied in detail problems of large plastic deformation. Due to uncertainties as to the slip systems at the time, terns, the final deformation gradient matrix can be obtained. Mathematically, the problem is reduced to finding the limit of F(dN as N — m while a — 0 in such a way that the Product Na = 0, a ftnite conslant designatzng the accunzulated amount of shear. It turns out that this limit is simply , where F, is the matrix whose elements are the coefficients of a in the F(a) matix. Application is made of the present treatment to a fcc crystal of Perlrzalloy compressed on the (110) plane and contrained to elongate zn the [112] directzon. In addition, it is shown that one rnay readily obtain from the general analyszs the well-known formulas relating elongation, amount of glide shear, and amount of lattice rotation for crystals deforming by single and double slip under tension. however, they were mainly concerned with proving that slip occurs on {111}(110) systems in fcc metals. The present general treatment arose out of recent investigations of magnetic anisotropy in cold-worked Fe-Ni alloys6 as well as some related problems in strength anisotropy of single crystals.' Detailed application of the general analysis will be provided in the case of a Permalloy single crystal compressed on the (110) plane and constrained to elongate only in the [112] direction. It will be seen that, as expected, the small strain approximation leads to significant errors after moderate straining. It will also be shown that the present general treatment yields the formulas of Mark et al. and of v. Goler and Sachs in those specific cases. GENERAL CONSIDERATIONS Consider a homogeneous deformation in which a material point initially at moves to both positions being referred to the same set of Cartesian axes. Such a deformation can be

Jan 1, 1967

By P. W. Bakarian

PREVIOUS determinations1,2,3 of the texture of magnesium and its alloys have shown only slight variations in the principal features of the structure. This investigation presents pole figures for magnesium and two of its alloys, in which differences of some magnitude will be shown. Schmid and Wassermann have considered the texture of magnesium in analogy with that of zinc-also a hexagonal lattice. Whereas the zinc texture is said to be derived primarily through basal slip and secondarily through {102}† twinning, the {102} twinning is considered impossible at the later stages of the rolling in magnesium because of its smaller axial ratio. This state of affairs appears reasonable for pure magnesium, but for alloy textures the situation is more complex. A double orientation similar to that found in zinc has been reported by J. D. Hanawalt,4 and the addition of small amounts of calcium to magnesium and its alloys to produce a double texture was noted by J. C. McDonald.5 The pole figures of this investigation, using an improved X-ray technique, show these various textures. The dependence of directional properties of rolled sheet on modifications of texture is well recognized. Correlation of physical properties with existing texture should be even more significant if quantitative valuations are attached to the various areas on the pole figures. Such a quantitative technique is developed herein. EXPERIMENTAL METHOD Two alloys were used, besides high-purity magnesium. In each case, the metal was extruded to bars 32 in. thick and hot-rolled to a thickness of about 0.075 in. It had been established previously that such a reduction was more than sufficient to eliminate the extruded structure. The rolling proceeded in the same direction as that of the extrusion, without any cross rolling. The temperatures corresponded to normal commercial practice, averaging around 600°F. A unique method of preparing the sample was developed in this work. Instead of using a flat specimen, a stem of the metal was cut out of the rolled sheet in the desired direction (rolling direction) as shown in Fig. 2. From a sheet of 0.075 to 0.065-in. gauge, the stem was cut to a square cross section with a jeweler's saw, rounded with a file and polishing paper to a diameter of 0.060 in, and evenly etched in dilute hydrochloric acid to a diameter of 0.038 ± 0.001 in. In this way, a cylinder of the original sheet with its axis in the desired direction was produced. By using this type of specimen for the pin-point pictures, any variance in the absorption of the diffracted beam at the various angular settings of the specimen, was completely eliminated.*

Jan 1, 1941

By W. E. Hall

The Darwin mining district of California, 160 miles north of Los Angeles, has yielded an estimated $45 million in lead, silver, zinc, and copper since 1875. The deposits are in silicated limestone of Penn-sylvanian and Permian age close to a stock of horn-blende-biotite-quartz monzonite. Steep left-lateral strike-slip faults that strike N 50o 70o E served as feeder channels for the ore solutions, and minor folds, favorable beds, and faults close to the feeder channels localized the orebodies. The orebodies are irregular, steep, pipe-like replacements that cut across bedding, bedding replacements, and fissure fillings in a calc-silicate country rock consisting mainly of wollastonite, grossularite-andradite garnet, idocrase, and diop-side. Near surface the ore consists of a soft, crumbly mass of cerussite, hemimorphite, jarosite, limonite, and some secondary copper minerals. The primary ore consists of galena, sphalerite, pyrite, pyrrhotite, and chalcopyrite with smaller amounts of andorite, matildite, scheelite, tetrahedrite, and a high silver-lead-bismuth-selenium-bearing galena. Preliminary X-ray fluorescence study of the primary ore has indicated unusually high concentrations of bismuth, selenium, and silver in galena from certain parts of the mine. One phase of this study concerns the distribution of these elements. Twenty-two samples of galena and thirty of sphal- erite have been purified and analyzed or are being analyzed. The selenium in 21 samples of galena ranges from 0.0045 to 2.11 pct; four contain more than 1 pct. These four samples were collected from the Essex orebody, where most of the selenium is concentrated. Three are from high-grade, near-surface ore, and one is from the 700 level. Four of the seven analyzed galena samples from the Thompson workings immediately north of the Essex orebody contain 0.29 to 0.74 pct Se; the others show less than 0.05 pct. Galena from the Defiance workings about 2000 ft south of the Essex has less than 0.03 pct Se. The selenium-bearing galena also contains concentrations of bismuth and silver. Preliminary spectographic analyses indicate as much as 3 pct Bi and more than 1 pct Ag in galena with high selenium content. Antimony is the only other minor element present in galena in amounts greater than 0.02 pct; most galena contains between 0.1 and 0.2 pct and the maximum is 0.72 pct. A second phase of the study is an attempt to determine the temperature range of the formation of the ore. The iron content of sphalerite deposited in equilibrium with pyrrhotite is being used as a geologic thermometer as it is a function of the temperature of deposition of the ore.' Preliminary results indicate temperature range from 430°C in replacement orebodies in the Defiance workings to 180°C in shallow fissure fillings.

Jan 1, 1960

[OFFICERS AND COMMITTEES J. W. Woomer, President A. B. Cummins, President-elect S. D. Michaelson, Past President C. E. Lawall, Eastern Regional Vice-President D. W. Scott, Central Regional Vice-President J. D. Forrester, Western Regional Vice-President R. B. Ladoo, Treasurer John Cameron-Fox, Secretary]

Jan 1, 1961

By F. W. Glaser

Iron-silicon alloys have had a great influence, in many ways, in modern industry. Silicon steels have been used almost exclusively for the construction of electrical machinery, but have also become an important material for structures which benefit by the high tensile and yield strengths of this material. In elaborating on the early history of work done on iron-silicon alloys, Greiner, Marsh and Stoughton1 went back as far as 1808, and pointed to Berzelius as the first producer of ferrosilicon. During the last twenty years some very important contributions in the form of studies of the iron-silicon system, as well as of the physical properties of these alloys, have been published. In consideration of the brit-tleness and poor machinability of alloys containing more than 5-6 pct Si, commercial interests have stimulated the research done on iron-silicon alloys, but mainly in the range of 0 to 6 pct Si. Yensen2 has shown the influence of im- purities, such as carbon, sulphur and phosphorus on the magnetic properties of 2, 4, and 6 pct Si alloys. Corson3 extended his investigation of physical properties to approximately 32 pct Si, and attributed the brittleness of the high-silicon material to the particular constitutional features of the alloys. Though mentioned earlier in the literature, the effect of nitrogen on crystal growth and magnetic permeability of iron-silicon alloys has been demonstrated recently by Morrill4 and Rostoker,5 respectively. In all the above quoted investigations, except that of Rostoker,5 sand or chill casting processes were employed for the preparation of samples. A major advantage of the powder metallurgy process is that it does away with time-consuming machining in many applications. The powder metal industry is always interested in pro- ducing a finished product employing a material which has presented considerably greater difficulties in manufacture by other methods. Steinit26 described preparation and testing of various Fe-Si alloys which were produced by the powder metallurgy process for the manufacture of armatures for small electric motors or generators. While Steinitz investigated alloys up to 6 pct Si, Rostoker5 went as high as 9 pct Si. The author7 previously discussed the diffusion of iron-silicon compacts containing 6 and 10 pct silicon. It is the object of this paper to study the progress of diffusion occurring in iron-silicon alloys by electrical resistivity measurements. All samples were prepared by pressing and sintering of compacts. Iron, silicon, and Fe-Si powders of various mesh size and compositions were employed to obtain test samples. It was assumed that complete diffusion had been obtained when the electrical resistivity reached a constant value characteristic for that particular alloy. To this end, electrical resistance measurements were used to follow the progress of alloying under various conditions.

Jan 1, 1950

By V. Carlson, J. A. Rowe, E. O. Bennett

This report concerns the microbial flora found throughout the surface facilities of six water-injection systems in Texas and Oklahoma. Each system is described in detail and water quality data are presented when available. Pseudomonads and sulfate-reducers (Desulfovibrio) were the predominate organisms found in the systems. The genera Sphaerotilus, Bacillus, Achromobacter, Minococcus, Clostridium, Flavobacteriurn, and Sarcina were also isolated in significant numbers. Molds, iron bacteria, sulfur bacteria and soil bacteria were found less frequently in the systems. It is concluded that a complex microbiological flora exists in oilfield water-injection systems. A tendency for waters to deteriorate in passage through solid bed filters is also noted. INTRODUCTION The injection of water into a subsurface formation is often accompanied by numerous problems, the origin and nature of which are only partially understood. During the past few years, investigators have indicated that bacteria are involved in some of these problems even though very little specific information is available concerning this matter. It is surprising that investigators working on microbiological problems in water injection have never made a complete study of the flora of these systems. The logical course to pursue is that of identifying the microbial flora and then examining each organism individually and in combination with others to determine if they produce a particular problem. The objective of the present investigation is to identify the microbial flora in water-injection systems. In addition. the quality of the water in most of the systems was determined and some observations made on the extent to which passage through surface-handling facilities affects water quality. The information derived from this study can serve as a foundation for additional basic investigations. EXPERIMENTAL PROCEDURE The microbial flora of six water-injection systems were studied at 50 sampling points. These systems are experiencing problems which may be attributable to micro-organisms, and two were being treated with antimicrobial agents at the time of this study. The samples were obtained by inoculating 1.0 ml of the water onto the different media at the sampling point. The media used for primary isolation included nutrient agar, Strokes agar, M-10-E medium and semi-solid thioglycollate medium. The cultures were returned to the laboratory, incubated at 30°C until growth occured; then, pure cultures were obtained from individual colonies. The pure cultures were identified according to standardbacteriologica1 procedures. Those organisms not fitting identification criteria exactly, even though exhibiting close similarities to known species, are listed as unidentified organisms. Membrane filters were used to make water-quality tests at intervals before and after collection of the microbiological samples. The filter discs were 47 mm in diameter with a 0.45-micron particle-size retention, and the tests were done under closed conditions at 20) psi. Membrane flow rates are plotted as increasing filtrate volumes, and the total solids contents arcs given when available. An effort was made to sample at each place in a system where a change in environment occurred. In this manner, the effect of passage through filters, tanks, hay sections, etc., on the microbial flora and water quality could be observed. RESULTS SYSTEM A This waterflood system (Fig. 1) is about four years old and daily pumps approx&ately 110,000 bbl of oil-free sour water from a 3,700-ft deep reef. The

By Charles H. Thurman

THE data given herein were first included in a paper read before San Francisco Section, A.I.M.E., in October 1940, and are applicable to conditions existing until the gold-dredging industry was temporarily suspended by Order L-208 of the War Production Board in the fall of 1942. There is no doubt that gold dredging will be resumed at some future time, and on a scale equivalent to that of the recent past, but the cost data given herein will be affected to some extent by conditions existing at that time, and by labor rates and prices for materials then prevailing. DRAGLINE VS. BUCKET-LINE Dragline dredges should not be considered for mining large areas of low-grade placer gravel. In such deposits a dragline operator with a 5-cu. yd. dragline would be doing well to maintain an operating cost of 6¢ per cu. yd., while a bucket-line dredge with buckets no smaller than 7 cu. ft. each could be operated for less than 4¢ per cu. yd. By operating costs, I mean only field charges such as direct operating and overhead, which includes pay roll, power or fuel, repairs and maintenance, insurance, land clearing and general expenses, but not including such items as prospecting, depletion, depreciation of equipment, State Franchise Tax, or Federal Income Tax. It is not to be inferred that a dragline dredge cannot make a profit in small low-grade deposits, where the extent of a given property is too small to justify the use of the more costly bucket-line dredge. Consider small deposits of 10¢ gravel. It is obvious that a deposit of 5,000,000 cu. yd. of such gravel could not be mined profitably with a new, 7-cu. ft. bucket-line dredge, costing upward of $250,000, and with operation costs at 4¢ per cu. yd., paying for land or royalties out of earnings; nor could a new combination 5-cu. yd. dragline and washing plant, worth upward of $125,000 and working at 6¢ per cu. yd., make a profit. These comparisons are made because the hourly capacity of both types of dredges is nearly the same. The bucket-line dredge averages 93 per cent operating time in 24 hr., as against about 80 per cent or less for the dragline dredge. These percentages of operating time are important factors in making a comparison of this kind. The bucket-line dredge has greater efficiency in cleaning soft bedrock, usually characteristic of dredging deposits, and a 7-cu. ft. bucket-line dredge can be efficient to a total digging depth of 80 ft. (65 ft. below water, 15 ft. above). A 5-cu. yd. dragline dredge is limited to about 30 ft. of efficient digging depth, 5 ft. of which could be above water. In small areas of gravel not over 30 ft. deep, a 5-cu. yd. dragline dredge could make a profit where a bucket-line dredge could not do so. As an example: compare the dredging of 3,000,000 cu. yd. of 10¢ gravel 28 ft. deep with bucket line, costing 4¢ per cu. yd., and with a 5-cu. yd. dragline dredge operating at 6¢ per cu. yd. Because of small yardage, new equipment is not warranted, therefore we can assume that the dredges of both types are used equipment, fully depreciated. Dismantling, mov-

Jan 1, 1945

By Edward C. J. Urban

ESSENTIAL requirements for ensuring safe working atmospheres in underground metal mines are planned systems of ventilation and provision for effective distribution of sufficient volumes of air by auxiliary equipment. Investigators1 in the past have suggested that quantities of air supply ranging from 300 to 800 cu. ft. of air per minute per man underground are necessary, to establish adequate general ventilation. These fundamentals are just as important as the use of water in dust suppression, control of blasting, the use of personal protective equipment, frequent inspection of working areas and education of employees in good dust-prevention practice. During the past two years the necessity for direct control of the underground working environment has become particularly urgent. The increase in production rate, with two-shift and three-shift operation replacing the single shift, has complicated the problem of dealing with excessive concentrations of blasting fumes and gases, carbon dioxide and monoxide, heat and oxygen deficiency. Although previous industrial hygiene programs had given consideration to these factors, they were generally regarded by mine operators as secondary to the control of dust hazards. Today, blasting smoke in a mining area means delayed production because the miner cannot return to his work place after a short interval. Smoke on the haulage level means a safety hazard and exposure of the underground employees to excessive concentrations of atmospheric contaminants. Exposure to the moderately excessive concentrations of carbon dioxide and monoxide found in metal mines has caused physiological disturbances and a decreased capacity for work; deficiency of oxygen, in some cases to as low as 3 per cent, has resulted in fatalities. During the past 10 years the Saranac Laboratory2 has been associated with the iron mines in the Upper Peninsula of Michigan and Wisconsin in developing and maintaining an industrial hygiene and silicosis-prevention program. At present, about 35 underground iron-ore mines are operated full time in this area. With two exceptions, all of them are ventilated by main fan installations installed either on surface or underground. ,The types of mining used in these mines may be classified either as sublevel cave, radial top slice or sublevel stope, and the procedures commonly employed in the ventilation of these three types are described in Figs. 1 and 2. A number of different methods of mine ventilation had been in use in this area when the study was started. Natural ventilation had proved unsatisfactory. The wide variation between low air volumes during the summer months and higher volumes during winter led to unpredictable conditions that were difficult to control. In the majority of the mines two separate shafts or openings to surface are utilized in

Jan 1, 1943

By Charles H. Thurman

The data given herein were first included in a paper read before San Francisco Section, A.I.M.E., in October 1940, and are applicable to conditions existing until the gold-dredging industry was temporarily suspended by Order L-208 of the War Production Board in the fall of 1942. There is no doubt that gold dredging will be resumed at some future time, and on a scale equivalent to that of the recent past, but the cost data given herein will be affected to some extent by conditions existing at that time, and by labor rates and prices for materials then prevailing. Dragline vs. Bucket-line Dragline dredges should not be considered for mining large areas of low-grade placer gravel. In such deposits a dragline operator with a 5-cu. yd. dragline would be doing well to maintain an operating cost of 6¢ per cu. yd., while a bucket-line dredge with buckets no smaller than 7 cu. ft. each could be operated for less than 4¢ per cu. yd. By operating costs, I mean only field charges such as direct operating and overhead, which includes pay roll, power or fuel, repairs and maintenance, insurance, land clearing and general expenses, but not including such items as prospecting, depletion, depreciation of equipment, State Franchise Tax, or Federal Income Tax. It is not to be inferred that a dragline dredge cannot make a profit in small low-grade deposits, where the extent of a given property is too small to justify the use of the more costly bucket-line dredge. Consider small deposits of 106 gravel. It is obvious that a deposit of 5,000,000 cu. yd. of such gravel could not be mined profitably with a new 7-cu. ft, bucket-line dredge, costing upward of $250,000, and in operation costs at 4¢ per cu. yd., paying for land or royalties out of earnings; nor could a new combination 5-cu. yd. dragline and washing plant, worth upward of $125,000 and working at 6¢ Per cu. yd., make a profit. These comparisons are made because the hourly capacity of both types of dredges is nearly the same. The bucket-line dredge averages 93 per cent operating time in 24 hr., as against about 80 per cent or less for the dragline dredge. These Percentages of operating time are important factors in making a comparison of this kind- The bucket-line dredge has greater efficiency in cleaning soft bedrock, usually characteristic of dredging deposits, and a 7-cu. ft. bucket-line dredge can be efficient to a total digging depth of 80 ft- (65 ft. below water, 15 ft. above). A 5-cu. Yd. dragline dredge is limited to about 30 ft. of efficient digging depth, 5 ft. of which could be above water. In small areas of gravel not over 30 ft. deep, a 5-cu. yd. dragline dredge could make a profit where a bucket-line dredge could not do so. As an example: compare the dredging of 3,000,000 cu. yd. of 10 gravel 28 ft. deep with bucket line, costin 4¢ Per cu. yd., and with a 5-cu. yd. dragline dredge operating at 6¢ Per cu. yd. Because of small yardage, new equipment is not warranted, therefore we can assume that the dredges of both types are used equip-ment, fully depreciated. Dismantling, mov-

Jan 1, 1946

By P. M. Robinson, M. B. Beaer

The heat of formation at 0°C of the intermetallic compound AgMg as a function of composition has been determined by tin-solution calorimetry. In this technique, the heat of formation is determined as the difference in the heat effects, at infinite dilution, on dissolution of the compound and a mixture of its components added alternately to liquid tin from a reference temperature, such as 0°C. The experimental technique and method of calculation have been described elsewhere.' The compound AgMg is a 0 electron compound with a CsCl structure.2 In nonstoichiometric compositions of the compound, atoms of the excess component are substituted on the AgMg lattice.3,4 The homogeneity range is approximately 43 to 52 at. pct Mg at room temperature and 38 to 65 at. pct Mg at 500Oc.2 Single-phase alloys outside the homogeneity range at room temperature can be prepared by quenching. The specimens used in this investigation contained from 39.0 to 54.8 at. pct Mg. They were made available, together with analyses, by the Research Laboratory of the General Electric Co. Metallographic examination of the quenched specimens containing 39.0 and 54.8 at. pct Mg gave no indication of the presence of a precipitated phase. The measured heats of formation are listed in Table I. In a calorimetric run, four to six additions each of a specimen and the corresponding mixture of the components were made. One or two calorimetric runs were carried out with each composition depending on the amount of samples available. The limits stated in the table represent the experimental precision and not the absolute accuracy of the measurements. Thermodynamic investigations of the compound AgMg have been made by Trzebiatowski and Terpilowski 5 and Kachi.6,7 The heats of formation at 500°C derived from the combined electromotive-force values of these investigators8 are more exothermic than the heats of formation at 0°C reported here, Fig. 1. The differences cannot be explained by the difference in temperature. This is shown by the published heat capacity of the stoichiometric compound, which obeys the Kopp-Neumann rule from 0" to 500oC.9 The heats of formation derived from equilibrium data are generally considered to be less accurate than those determined by calorimetric methods because small errors in the change of the equilibrium values with temperature have a large effect on the calculated value of the heat of formation. The heats of formation derived from the combined published electromotive-force values reach a maximum at 56.0 at. pct Mg.8 Kachi reports a maximum of -4.7 kcal per g-atom at the stoichiometric composition.7 The present results also indicate a maximum at or near the stoichio metric composition, Fig. 1.

Jan 1, 1964

By Nicholas J. Grant, H. C. Chang

EMPHASIS on the importance of grain boundary sliding as a mode of deformation at elevated temperatures has been presented elsewhere.' The extent to which boundary sliding occurs under certain creep conditions has been determined for high purity aluminum2,3 and aluminum alloys.' In general, boundary deformation extends quite far into the adjoining grains, resulting in a rather wide band of deformation along the grain boundaries. This has led some investigators to doubt that grain boundary sliding consists of shearing along the boundary surface; it has been contended instead that boundary sliding is a result of concentrated slip in the neighborhood of the deformed grain boundaries."' The purpose of this note is to present evidence that grain boundary sliding takes place by a bulk shear process. gh purity polycrystalline aluminum, wherever and whenever there is grain boundary sliding, boundary migration and concentrated deformation along both sides of the deformed boundary are associated with it. Therefore, it has been rather difficult to obtain unambiguous evidence of boundary sliding. On the other hand, it would be expected in certain alloys of aluminum, when tested under particular conditions. that the severity of the interfering effects of boundary migration and concentrated deformation along the deformed boundary might be reduced. Current research in this laboratory' shows that a 20 pct Zn-A1 solid solution alloy fractures in an intercrystalline manner during creep at 500°F with a small total elongation of not more than about 14 pct, and that it does so with extremely limited grain boundary migration. Accordingly, this alloy was chosen for the present study. The experimental procedure was similar to that used in previous work and has been presented elsewhere.' Two parallel flat surfaces were milled from a round specimen. The gage portion of the specimen was 1 in. x .09 in. x .17 in. The specimens, which were annealed at 1030°F for 24 hr, showed two to four grains across the width of the test bar. The specimens were subjected to creep at 500°F at stresses of 2000 to 5000 psi. The rupture times were from 10 sec to 3 hr. Appreciable boundary sliding occurred in all the tests. A typical example of boundary sliding is shown in Fig. 1. (The tension axis is vertical in the micrographs.) Sliding has taken place along the boundary between grains A and B unaccompanied by important boundary migration. The sharpness of the offset at the triple point can be clearly seen in grain C. As a result of sliding along the boundary between grains A and B, a new boundary surface between grains A and C was created. The offset at the triple point is about 0.07 mm. This boundary sliding is accompanied by the formation of rather sharp kinking bands and, of course, slip on several slip systems in grain C. This type of deformation has been referred to as fold formation' in previous work with high purity aluminum.

Jan 1, 1957

By E. C. Houston, H. S. Rankin

Olivine is considered a valuable potential source of metallic magnesium in the chloride electrolytic process. Treatment of olivine with hydrochloric acid can be carried out under conditions that prevent the formation of gelatinous silicic acid, which would inhibit the reaction by forming a covering around the ore particles. It is our hypothesis that the formation of gelatinous silicic acid is avoided by the dehydrating action of the magnesium chloride, which takes up water to form a series of hydrates and leaves the silicic acid in granular condition. Leaching serves to separate the magnesium chloride from the silicic acid and unattacked ore, which can be treated with fresh acid in successive steps until extraction is essentially complete. The leach solution can be i purified by the addition of caustic magnesia and the precipitated impurities removed by decantation or filtration. The resulting magnesium chloride contains: Fe, less than 0.001 Per cent; Mn, 0.012 ; per cent; Ni, none. Chlorination† of olivine has two major defects: (1) a large part of the chlorine goes to form silicon and iron chlorides, (2) the anhydrous magnesium chloride is held by absorption in the finely divided silica. SupplY It is well known that olivine, a silicate of magnesium and iron, is present in large amounts in several localities in the United States. The largest deposits occur in the southeast and on the Pacific Coast.' The southeastern deposits were surveyed by Lewis2 in 1896 and by Hunter3 in the period 1933-1940 It is estimated that high-grade olivine, conveniently located with regard to transportation and power, is present in excess of 230 million tons above local drainage level in western North Carolina and north Georgia. Previous Technical Work In the past few years, olivine has attracted considerable attention in the refractory field. In the chemical processing industry, however, it has long been thought axiomatic that the cost of processing silicates is an economic barrier to their utilization, especially where more reactive minerals, such as carbonates, oxides or salts, are available. This attitude is not entirely justified in connection with olivine. unlike most silicates, olivine is readily attacked by acids, chlorine and other reagents without resorting to fusion to break down the silicate structure. Olivine has been suggested as a raw material for making epsom salt by roasting with pyrites.4 Other processes include acid treatment6 and digestion with calcium chloride.6 Recently a small plant was built near Webster, N. C., for making epsom salt from olivine by treatment with sulphuric acid. This plant is now in commercial operation, producing approximately two tons per day. Olivine is also receiving attention from the fertilizer industry as an inexpensive source of soluble magnesium.

Jan 1, 1942

By H. S. Rankin, E. C. Houston

Olivine is considered a valuable potential source of metallic magnesium in the chloride electrolytic process. Treatment of olivine with hydrochloric acid can be carried out under conditions that prevent the formation of gelatinous silicic acid, which would inhibit the reaction by forming a covering around the ore particles. It is our hypothesis that the formation of gelatinous silicic acid is avoided by the dehydrating action of the magnesium chloride, which takes up water to form a series of hydrates and leaves the silicic acid in granular condition. Leaching serves to separate the magnesium chloride from the silicic acid and unattacked ore, which can be treated with fresh acid in successive steps until extraction is essentially complete. The leach solution can be i purified by the addition of caustic magnesia and the precipitated impurities removed by decantation or filtration. The resulting magnesium chloride contains: Fe, less than 0.001 Per cent; Mn, 0.012 ; per cent; Ni, none. Chlorination† of olivine has two major defects: (1) a large part of the chlorine goes to form silicon and iron chlorides, (2) the anhydrous magnesium chloride is held by absorption in the finely divided silica. SupplY It is well known that olivine, a silicate of magnesium and iron, is present in large amounts in several localities in the United States. The largest deposits occur in the southeast and on the Pacific Coast.' The southeastern deposits were surveyed by Lewis2 in 1896 and by Hunter3 in the period 1933-1940 It is estimated that high-grade olivine, conveniently located with regard to transportation and power, is present in excess of 230 million tons above local drainage level in western North Carolina and north Georgia. Previous Technical Work In the past few years, olivine has attracted considerable attention in the refractory field. In the chemical processing industry, however, it has long been thought axiomatic that the cost of processing silicates is an economic barrier to their utilization, especially where more reactive minerals, such as carbonates, oxides or salts, are available. This attitude is not entirely justified in connection with olivine. unlike most silicates, olivine is readily attacked by acids, chlorine and other reagents without resorting to fusion to break down the silicate structure. Olivine has been suggested as a raw material for making epsom salt by roasting with pyrites.4 Other processes include acid treatment6 and digestion with calcium chloride.6 Recently a small plant was built near Webster, N. C., for making epsom salt from olivine by treatment with sulphuric acid. This plant is now in commercial operation, producing approximately two tons per day. Olivine is also receiving attention from the fertilizer industry as an inexpensive source of soluble magnesium.

Jan 1, 1942

By J. L. Wyatt

The electrical resistance of titanium as a function of purity and temperature was measured from —325" to 2800°F. Two points of inflection in the data plots were found, and an increase in resistance with increase in temperature above the transformation was observed. IN 1948 Greiner and Ellis' reported on the thermal and electrical properties of titanium, including a study of the effect of temperature on the electrical resistance. The materials used in the research were obtained from the United States Bureau of Mines, where powder metallurgy techniques were being employed for consolidation purposes. These authors found a decreasing resistance with increasing temperature above 1625°F, and concluded that the latter temperature represented the point of transformation from hexagonal close-packed to body-centered cubic structure on heating. They reported a specific resistivity value of 55x10-6 ohm-cm, and showed that dp/dt was positive, decreasing with increasing temperature above ambient. In one of his early texts Hume-Rothery2 mentioned that certain metals, including titanium, exhibited minima in plots of resistance vs temperature. He indicated that this occurred at 212°F. Reporting many years ago on the effect of temperature on the electrical resistance of zirconium, de Boer and Fast³ showed that phase transformation in this sister metal occurred over a range of temperatures, the extent varying with its purity. Graphs of resistance vs temperature, showing transformation effects over a range of several hundred degrees, were presented for relatively impure metal. In the latter case the resistance decreased with increasing temperature throughout the region of partial phase transformation, thereafter increasing as predicted for normal metals by modern theories of the solid state. Since the quality of metal used in previous studies was substandard compared with iodide-type titanium, and in view of the fact that there seemed to be some rather unusual observations in the reported literature, it was decided to undertake a more extensive study of the thermal effects using high purity titanium, and to include temperature ranges beyond those which had been studied in order to ascertain whether or not the characteristics of zirconium would be duplicated. Experimental Materials Materials used in this investigation included commercial quality titanium (Ti-75A produced by the Titanium Metals Corp. of America) in the form of ¼ in. diam rods, and high purity iodide titanium (produced by the Foote Mineral Co.) in the form of crystal bar of irregular size. The latter was cold-swaged into rod form and machined to the desired size. Lead wires to all specimens consisted of 18 gage Ti-75A. Chemical analyses of the stock materials are given in Table I. Experimental Procedures Specimens 2½ to 3 in. in length were prepared with diameters nominally 0.15 to 0.20 in. Two holes,

Jan 1, 1954

Engineering Experiment Station, University of Arkansas, Fayetteville, Ark. W N Gladson, Dean and Director The Engineering Experiment Station has published a number of bulletins, only a few of which are of particular interest in this directory: Bulletin 3, Arkansas coal or oil for domestic heating; 4, Road materials of Arkansas, 5, A study of Arkansas coals, 6, Portland cement materials in Arkansas; 11, Natural gas versus Arkansas coal for house heating. Circular 3, Short cuts to combustion efficiency with natural gas Address the Engineering Experiment Station for any of these publications.

Jan 1, 1933

By Joseph G. Franzone, Alan K. Kuhn, Jerry J. Oliver

INTRODUCTION Tailings produced by many uranium milling operations consist of fine sands with some silt and slimes. When fine sandy materials are placed by hydraulic methods behind an impermeable dam, they are deposited in a loose state and some portion remains saturated. These saturated, loose fine sands are highly susceptible to liquefaction, the phenomenon in which a monotonic or cyclic disturbance causes excess pore water pressure and, concurrently, a total loss of strength in the sand. Once liquefied, the fine sand flows, causing collapse of supported structures. The potential for such a liquefaction failure was identified and successfully treated at Sohio Western Mining Company's L-Bar Uranium Operations near Seboyeta, New Mexico. The Sohio tailings dam is an earthfill starter dam with tailings raises constructed upstream of the starter dam crest. Beginning with the start of mill operations in 1976, tailings were deposited hydraulically in the impoundment behind the starter dam. As the tailings level approached the crest of the starter dam, beached tailings were excavated and placed in lifts by a dozer to form two raises above the crest of the starter dam. As the impoundment has gradually filled with tailings, the decant water pond level has risen. Correspondingly, the water level in the tailings dam, upstream of the chimney drain, has also risen. By August 1980, the phreatic surface was above the crest of the starter dam and stability analyses indicated that the tailings dam did not have the safety factor of 1.0, under static loading with liquefaction, implicit in the 1980 New Mexico State Engineer's requirements. The problem related directly to the high phreatic surface in the loose sandy tailings behind a dam structure that was not capable of resisting failure which might develop in liquefied tailings. Two methods of stability enhancement were considered for the Sohio dam. One method was reinforcement of the dam using a clay buttress placed against the downstream slope. The other method was installation of horizontal drains through the starter dam to drain decant water and lower the phreatic surface. Either method would increase the stability of the dam. However, the comparative analysis of the methods indicated that horizontal drainage offered significant advantages in cost and could provide a more direct remedy of the source of the problem, the saturated section of tailings. In addition, a drainage system would be more amenable to modification and more quickly installed than a buttress. On the other hand, a clay buttress could be more directly controlled, would involve routine earthwork, and has successful precedent. Therefore, the buttress approach would involve less risk. After weighing the advantages of both methods and the costs versus risks, the horizontal drainage approach was selected. Consequently, a horizontal pipe drain system was recommended and designed (D’Appolonia, 1981a). This paper describes the design and installation of the drainage system and its initial performance. PIPE DRAIN SYSTEM DESIGN Design Criteria and Parameters To increase the stability of the dam to acceptable levels, the drainage system had to lower the phreatic surface behind the dam to the level calculated to give a minimum safety factor of 1.0 under conditions of liquefaction. In addition, the system needed to be essentially self-operating, easy to maintain, and economical. To accomplish these objectives, the following design criteria were established: 1. The pipe drains must lower the phreatic surface at least 4.3 m (14 ft.) at the midpoint along the dam (the deepest section) and by proportionately smaller amounts toward the abutments to satisfy safety factor requirements.

Jan 1, 1982

By F. H. Wilson, E. W. Palmer

Brass mills are familiar with a recurring problem which reveals itself during deformation of annealed metal as an opening up of cracks which are suggestive of a grain boundary pattern. A typical example is seen in Fig 1, which shows part of the convex surface of a cartridge brass disc which was slightly dished by the punching operation. Another illustration of the same type of defect was found in the surface of a finished cartridge case which split open on firing. These cracks, shown in Fig 2, were away from the split but indicate the presence of the type of weakness which permitted the splitting. Usually the weakness consists of separate cracks, the lengths of which are of the same order of magnitude as the grain size prior to the final anneal. Their typical appearance in a polished section is shown in Fig 3. This structure was found below the surface of a fractured tensile-specimen that had been cut from a large cartridge blank showing the defects on its convex surface. While the pattern formed by these cracks suggests a grain boundary origin, the cracks bear no relation to the currently existing grain structure. Thus Fig 3 shows that the grains have grown up to cracks already present. Since it is doubtful that the cracks were present in the metal before the anneal, they must have formed during the anneal but prior to recrystallization. The term "fire-cracking" has been generally applied to cracking that occurs during an anneal under the influence of internal stress, but refers more specifically to obvious macroscopic cracking attributable to the presence of a low melting phase, usually lead. Since the type of cracking described above may occur when the lead content is very low, we have considered it a somewhat different phenomenon, and have called it "inter-granular parting." While most examples involve cartridge brass, intergranular parting has also been observed in the fabrication of large seamless tubes from discs of both 85/15 red brass and 70/30 cupro-nickel. Similar cracking in nickel silver was investigated by Jones and Whitehead,1 who showed that it could occur during healing or cooling. That which occurred on heating they called "fire cracking," and they suggested that, a transformation at about 320°C might account for its occurrence. It was felt in this laboratory that the observed parting was probably one aspect of the general observation, first made by Rosenhain and Archbutt,2 that a tendency toward intergranular fracture under tensile stress increases with increasing temperature and decreasing rate of strain. In this case the stress involved would be internal. The rate of strain would be exceedingly slow, a localized internal creep. Reference to the literature uncovered no efforts to extend the observations of Rosenhain to conditions involving only internal stress. Accordingly, an exploratory research, sufficient in detail to satisfy us as to the probable truth of this explanation of the observed intergranular parting, was undertaken. The internal stresses present during the early stages of an anneal (prior to recrystallization) would be from two sources: residual stresses developed during deformation, and thermal stresses caused by uneven heating and expansion. Thermal stresses would vary widely according to shape, size and manner of heating, and can be considered as supplementary to residual stresses. It seemed necessary to determine the stress and temperature conditions under which parting would occur in a relatively short time, and then to establish whether or not in- ternal stresses may persist to an extent adequate to cause parting at such temperatures. (Jones and Whitehead1 and Moore and Beckinsale3 made tests which indicated that stress relief required an appreciable time.) During the course of the research the desirability of studying the effect of grain size became apparent, and this factor is one of the major variables of our investigation. Intergranular Parting under Tension at Elevated Temperatures In early experiments, hard 70/30 brass tensile specimens, with unknown residual stresses, held for 10 min. at applied stresses from 30,000 to 40,000 psi at temperatures from 300 to 350°C, showed no macroscopically visible cracks when unloaded and cooled, but cracks very similar in appearance to those observed in commercial practice were revealed in these specimens by pulling them to fracture. Annealed specimens, on the other hand, showed no cracks in such experiments, apparently because they deformed plastically at stresses below those necessary to cause parting in a reasonable time. Hence, the specimens used for this study were first strengthened by cold tensile elongation, giving them an unavoidable residual stress pattern (determined largely by grain size and orientation) which would, however, correspond in direction, at least, to the stress pattern obtained under stress at temperature. A variation in these residual stresses might be expected with variations in grain size, but following any given anneal the reproducibility of stress pattern should be as good as that of the grain size measurement. With no information as to the effect of the amount of prior deformation, tests were conducted on specimens given the same cold elongation, using a stress applied at temperature which was a constant proportion of the stress required to produce this elongation. In

Jan 1, 1950