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By Blake M. Loring

Recently the S-curves for 30 to 40 alloy steels have been pub1ished.l.2 These steels show individual characteristics, which make each additional S-curve of great interest. There are important differences in the mechanism of the transformation process, the type of transformation product, and the influence of alloying elements. Steels of the same general alloy composition may possess such variable characteristics with regard to the length of time of isothermal transformation that an S-curve for each heat of steel may be desirable. In the present work the isothermal transformation of a steel having the following composition and grain size was studied: 0.29 per cent carbon; 0.21 manganese; 0.026 phosphorus; 0.017 sulphur; 0.056 silicon; 1.45 chromium; 3.25 nickel; austenite grain size, 7. The progress of the isothermal transformation was followed by several different methods, including hardness measurements, dilatometer observations, and examinations of the microstructure. It was not expected that these methods would be in precise agreement.3 Accordingly, the changes in microstructure and hardness were given additional weight in determining the initial stage of transformation and the dilatometer measurements were relied on in estimating the completion of the process. The beginning of transformation was indicated by I per cent of transformation products and the end of transformation by about 99 per cent. The reproducibility of the various methods in the ranges of highest sensitivity was checked by several observations at each time and temperature. The steel used in this investigation was annealed for 24 hr, at 925°C. Macroetching revealed no serious segregation. The specimens for the metallographic and hardness investigations were 1¼ by ½ by ? in. Wires attached to the samples were used for handling. instead of tongs, to avoid chilling. The specimens were heated at 845°C. for ½ hr. and quenched into constant-temperature salt baths held at various temperatures from 205' to 650°C. The time required to transfer the specimens was found to be about 1½ sec., and nearly 5 sec. more was required for them to come to the temperature of the bath. After various lengths of time in the constant-temperature baths the specimens were quenched in water at 2I°C. The specimens were cut in half for the metallographic examination, to provide a surface free from decarburization. To determine the beginning and end of the transformation the percentage of transformation determined metallographically and by hardness was plotted against time. A planimeter was used in the metallographic determination. The hardness and metallographic determinations made at room temperature were confirmed reasonably well by dilatometer measurements taken during the process of transformation at the various temperatures. Two dilatometers were used to

Jan 1, 1942

By Blake M. Loring

Recently the S-curves for 30 to 40 alloy steels have been pub1ished.l.2 These steels show individual characteristics, which make each additional S-curve of great interest. There are important differences in the mechanism of the transformation process, the type of transformation product, and the influence of alloying elements. Steels of the same general alloy composition may possess such variable characteristics with regard to the length of time of isothermal transformation that an S-curve for each heat of steel may be desirable. In the present work the isothermal transformation of a steel having the following composition and grain size was studied: 0.29 per cent carbon; 0.21 manganese; 0.026 phosphorus; 0.017 sulphur; 0.056 silicon; 1.45 chromium; 3.25 nickel; austenite grain size, 7. The progress of the isothermal transformation was followed by several different methods, including hardness measurements, dilatometer observations, and examinations of the microstructure. It was not expected that these methods would be in precise agreement.3 Accordingly, the changes in microstructure and hardness were given additional weight in determining the initial stage of transformation and the dilatometer measurements were relied on in estimating the completion of the process. The beginning of transformation was indicated by I per cent of transformation products and the end of transformation by about 99 per cent. The reproducibility of the various methods in the ranges of highest sensitivity was checked by several observations at each time and temperature. The steel used in this investigation was annealed for 24 hr, at 925°C. Macroetching revealed no serious segregation. The specimens for the metallographic and hardness investigations were 1¼ by ½ by ? in. Wires attached to the samples were used for handling. instead of tongs, to avoid chilling. The specimens were heated at 845°C. for ½ hr. and quenched into constant-temperature salt baths held at various temperatures from 205' to 650°C. The time required to transfer the specimens was found to be about 1½ sec., and nearly 5 sec. more was required for them to come to the temperature of the bath. After various lengths of time in the constant-temperature baths the specimens were quenched in water at 2I°C. The specimens were cut in half for the metallographic examination, to provide a surface free from decarburization. To determine the beginning and end of the transformation the percentage of transformation determined metallographically and by hardness was plotted against time. A planimeter was used in the metallographic determination. The hardness and metallographic determinations made at room temperature were confirmed reasonably well by dilatometer measurements taken during the process of transformation at the various temperatures. Two dilatometers were used to

Jan 1, 1942

NEW MEMBERS The following list comprises the names of those persons who became members during the period of Oct. 10, 1917, to Nov. 10, 1917. ABEY, HIROSHI Min. & Met. Engr.. Mitsubishi Co., Tokio, Japan. BARNETT, WALTER W., Assayer & Chem., Seoul Min. Co., Tul Mi Chung, Chosen, Korea. BATES, WILLIAM S Min. Engr., Anaconda Copper Min. Co., Butte, Mont. BENDELARI, ARTHUR E Mgr. of Mines, Eagle Picher Lead Co., Picher, Okla. BERKEFELD, PAUL C Mining Engr., 2901 Piedmont Ave., Berkeley, Cal. BILLINGS, It. S Mgr., Sheep Trail Pyramid Mines Co., Kingman, Ariz. BILLWILLER, C. J., JR Mgr. Nitrate Agencies, Ltd., Iquique, Chile. BLUM, EDWARD H., Engr., Asst., to Pres., The Atlantic Refining Co., 3144 Passyunk Ave., Philadelphia, Pa. BORDEN, GAIL, Vice-Pres. & Treas., American Safety Explosive Co., 50 East 42d St., New York, N. Y. BRESLIN, THOMAS F Min. Engr., Cebu, Cebu, P. I. BURKE, JAMES M., Efficiency Engr., Calumet & Hecla Mining Co., Calumet, Mich. CANTON, WILLIAM R Met., Nevada Consolidated Copper Co., McGill, Nev.

Jan 12, 1917

By Robert Linton

Tats paper deals with work that has been carried on for over three years by the management and staff of the North Butte Mining Co. in an effort to standardize mining methods, to eliminate lost motion in the operating organization, and to improve the efficiency of the individual workman. Naturally, much of this work is an adaptation of ideas suggested by other mining operations or by practice in manufacturing and other productive industries. The studies, however, have developed some features that it is believed are new. Both new and adapted features are given in detail. The normal capacity of the North Butte mines is about 2000 tons per. day. Hoisting is done principally through the Granite Mountain shaft, which is 3740 ft. (1140 m.) deep and equipped with an electric hoist capable of hoisting 200 tons per hour from a depth of 4000 ft. (1219 m.). The lowest stoping level is the 3400-ft. The Granite Mountain shaft is connected with the Speculator shaft, about 850 ft. (259 m.) south, and with the Gem, about 1150 ft. (350 m.) east, which are 3000 (914.4 m.) and 2000 ft. (609.6 m.). deep, respectively. These two shafts are equipped with reversible fans for ventilation. The Speculator has a; steam hoist with a capacity of 150 tons per hour from the 3000-ft. level. The Granite Mountain shaft is concrete lined from the 1600-ft. to below the 3000-ft. level and gunited with cement from the 1600-ft. level to the surface. Sections of the Speculator shaft with a total combined length of 500 ft. (152.4 m.) are also concreted. There are approximately 36 miles (58 km.) of drifts and crosscuts in the mine, in which there are 15.2 miles (24.5 km.) of track. At normal capacity, about 75 stoping faces, 20 raises, and 30 sill development faces are working. Stoping is usually by the standard square-set method. Where the size of the oreshoot and character of. the walls permit, the rill (or cut-and-fill) system is employed.

Jan 8, 1920

By Guy T. McBride

The 1970's have been considered a pivotal decade for the mineral industries. Undergraduate mineral engineering education is relevant to this theme; in particular, there are three interrelated problems: 1) The goals, and therefore the course requirements, of the under- graduate curricula are in question; 2) The unit costs of undergraduate education appear excessive to those who bear the charges; and 3) The present social ambience detrimentally affects the work attitudes and career motivation of both prospective and in-course students. Any discussion of these problems must bottom upon recognized national needs for minerals and a technologically sound minerals industry staffed by competent engineers, particularly at the exploration, development and production levels. Fortunately, these needs have been recognized and to some extent quantified, and they are even now under further intensive study by some of the most competent members of society and the profession.

Jan 3, 1973

By G. F. Sager, B. J. Nelson

Magnesium alloy forgings offer higher and more uniform mechanical properties than heat treated magnesium alloy castings and are used principally for light weight parts that may be stressed in fatigue or subject to impact loads, or where space limitations do not permit the use of the necessarily bulkier magnesium alloy castings. Compositions and Properties of Magnesium Forging Alloys The nominal compositions and properties of the principal commercial magnesium forging alloys are given in Table 1. The two that are most commonly used are AMC58S and AM65S. The former has the highest strength and is, therefore, employed for heavy duty applications. However, the forging of this alloy and of AMC57S ordinarily requires the use of slow speed hydraulic presses. The forging alloy AM65S combines good, although somewhat lower, mechanical properties with excellent forging characteristics. This alloy can be forged on hammers, upsetters, rolls, and mechanical presses, as well as on hydraulic presses. These characteristics make it particularly useful for many applications. Manganese Segregation in AM65S The alloy AM65S contains 5 pct tin, 3 1/2 pct aluminum, and a small amount of manganese, the manganese being added to insure satisfactory resistance to corrosion. Some manganese segregation was occasionally encountered in this alloy. In many applications, such segregation is of no particular consequence, but when forgings are acid dipped prior to the application of various chemical coatings, the segregated particles of manganese constituent may stand in relief. The resultant surface roughening may impair appearance and is undesirable in forgings employed in the textile industry, where very smooth surfaces are essential to avoid the catching and breaking of fine threads. An effort was therefore made to determine the cause of manganese segregation in this alloy and find a method for its elimination. Radiographic examination of a series of cast ingot sections and extruded samples of the AM65S type having manganese contents ranging from zero to about 0.7 pct revealed no appreciable manganese segregation when the manganese was less than about 0.4 pct. A small amount was observed with a manganese content of 0.46 pct and considerable with manganese contents of 0.6 pct or higher. Attempts to eliminate the manganese segregation by variations in the melting and casting conditions were unsuccessful and it soon became apparent that the trouble was attributable to the fact that the manganese contents that caused segregation were in excess of the amounts that would remain in solution at temperatures encountered in the melting and/or casting operations prior to the complete solidification of the ingots. A knowledge of the amounts of manganese that would be soluble in molten alloys of the AM65S type at various temperatures was obviously desirable. The liquid solubilities of manganese in binary Mg-Mn alloys and in Mg-Al-Mn and Mg-Zn-Mn alloys have been determined by Tiner and others,* but no data on the solubility of manganese in molten magnesium alloys containing tin and aluminum are given in the literature. For this reason, an investigation to determine the solubility of manganese in a magnesium alloy containing 5 pct tin and 3.5 pct aluminum (AM65S type) was undertaken.

Jan 1, 1950

By R. E. Zimmerman, Michael Sokaski, E. R. Palowitch, M. R. Geer, H. F. Yancey, W. Deurbrouck, S. C. Sun

PART 1: DENSE MEDIUM SEPARATION by M. SOKASKI, M. R. GEER and H. F. YANCEY INTRODUCTION In the early days of coarse-coal treatment by the dense-medium process in Europe, loess was one of the materials used for medium solids. Newly developed cyclones were used to reclaim and thicken :he loess suspension. On occasion, when the cyclone thickener at the Maurits mine in the Netherlands plugged, the overflow was found to be filled with clean coal free of impurity. From this observation stemmed the concept of the cyclone as a cleaning device. Development took place during World War 11 in the laboratory of the Dutch State Mines under the direction of M. G. Driessen. A 15-ton-per-hour pilot plant was in operation in the laboratory by 1945, when, in September, Driessen made the first public disclosure of the process in a paper presented to the Institute of Fuel in London. In April 1945,2 operation of this pilot plant was observed by one of the authors, who was so impressed with the potential of the cyclone process that, on his return to the United States, a research program on the cyclone was started by the Bureau of Mines immediately. This work led to an AIME paper in 1946 in which the first details of cyclone operation on American coals was made available. Commercial adoption of the cyclone proceeded in Europe, initially in the Netherlands, Germany, and France. In the United States, where coal is less costly and most of the coals are so easy to clean they respond well to simpler cleaning systems, the first cyclone plant was not built until 1961.3 As of the spring of 1967, a total of 31 commercial cy- clone plants had been installed or were under construction in the United States; this does not include eight Dynawhirlpool plants. The capacity of the cyclone plants ranges from 50 to 1300 tons per hour and totals 6600. THEORY Before discussing the theory of a dense medium cyclone, a brief description may be helpful to some readers who are not familiar with its basic operation. In a typical dense-medium cyclone, illustrated in figure 10-1, the mixture of medium and raw coal enters tangentially near the top of the cylindrical section, thus forming a strong vertical flow. The refuse moves along the wall of the cyclone and is discharged through the under- flow orifice. The washed coal moves toward the longitudinal axis of the cyclone and passes through the vortex finder to the overflow chamber.

Jan 1, 1968

By Donald R. Mason, Daniel F. O’Kane

DIFFERENTIAL thermal analysis measurements have been used to determine the ZnTe-In2Te3 pseudo-binary phase diagram. Since ZnTe and In2Te3 are both semiconductors, any intermediate compounds formed in this system should also be semiconductors according to the rules delineated by Mooser and pearson.1 The compound ZnIn2Te4 was first reported by Hahn et a1.2 and it has subsequently been prepared and characterized electrically by several workers. A brief summary of this work has been presented by Mason and O'Kane3 who also report that the compound appeared to melt peritec-tically. The Zn-Te phase diagram has been investigated by Kulwicki4 who found that ZnTe has a melting point of 1290° + 2°C. The In-Te diagram published in Hansen' contains significant errors, and has been redetermined by Grochowski et al.6 However, both sources agree that the melting point of In2Te3 is 667°C, with the latter authors reporting the previously observed solid-solid transformation occurring at about 550°C. Hahn ef al.2 reported lattice-parameter values for the ZnTe-In2Te3 system. The 67 mole pct ZnTe-33 mole pct In2Te3 composition was the only

Jan 1, 1965

By D. L. Harris

Chemical upgrading testwork on Stillwater chomite concentrate shows that reduction roasting and leaching (sulfuric acid and/or ferric sulfate) can provide upgraded feed to ferrochromium furnaces for making upgraded ferrochromium. Two-stage reduction smelting in electric arc furnaces can make upgraded ferrochromium also. Sulfating roasts and leaching (acid and/or water) results using gaseous SO2—SO3 or solid ammonium sulfate were not satisfactory for upgrading chromite. Chemical upgrading of domestic Stillwater (Montana) chromite has been considered at various times1'2' to provide a material with higher chromium to iron ratio, more like foreign high grade ore, for the production of standard grade ferrochromium. Physical bene-ficiation of Stillwater chromite ore can provide a ratio, Cr/Fe, no higher than the ratio in the chromite mineral, which is approximately 1.7. For comparison the analyses of a foreign metallurgical grade chromite ore, a high grade electrostatic process Stillwater concentrate and a stockpile concentrate from the Stillwater ore are noted in Table I. Single stage smelting of the high grade foreign ore with a ratio, Cr/Fe, of 3.37 can provide a ferrochromium of over 65 pct chromium content. Similar smelting of the Stillwater concentrate with a ratio of 1.7 can provide a ferrochromium of 50 to 55 pct chromium content.' It has been demonstrated on a pilot plant scale that chemically upgraded Stillwater chromite having a ratio, Cr/Fe, of 3.0 also can be smelted to provide a ferrochromium containing over 65 pct chromium.5 Two major considerations for an upgrading process are the ratio, Cr/Fe, in the product and the chromium recovery. A minimum ratio, Cr/Fe, of 3.0 is required in the product but for economic reasons a higher ratio is desirable for blending purposes as long as the recovery is satisfactory. Data from different kinds of chemical upgrading testwork are presented in this report. Some of the testwork was conducted to make a product which was not intended to be feed for ferrochrome smelting, but the data are considered pertinent to chemical upgrading. Results reported by others are noted also for comparison. WORK REPORTED BY OTHERS The U. S. Bureau of Mines has contributed much to the literature on chemical upgrading of chromite and a summary is given in Table 11. Reduction roasting and acid leaching were reported by Boericke and Bangert,1 by Lloyd and Gorst, et a1,2 and by Boericke3 to demonstrate very good success on a laboratory scale and only fair success in pilot plant tests for Stillwater (Benbow) concentrates. This work included basic studies on thermodynamic properties to explain the reduction and leaching results. Canadian workers, Downes and Morgan,9 have reported experimental success in upgrading a low grade Bird River chromite, similar to Stillwater material by the same process. Russian workers1' have noted results on reduction tests of high grade chromite which may apply to low grade ores and which help in understanding the reduction process. The use of flux additives in the reduction process was an innovation of their work. The Canadian workers9 also reported other kinds of experiments, including autoclave leaching of the chromite in a limited amount of sulfuric acid. An insoluble basic chromium sulfate was formed and separated from soluble iron, magnesium, and aluminum sulfates to provide an upgraded material. U.S. Bureau of Mines workers, Lloyd, Rosenbzum, Home and Davis,' reported on autoclave leaching of Benbow concentrates as a means of obtaining chromium alum for electrolysis. This last work required selective crystallization of alum to separate iron from chromium. Chloridizing roasting of Benbow and Mouat concentrates has been reported by Hunter and Banning6 and others8 to show that the iron chloride could be volatilized from the chromite to provide upgraded residues.

Jan 1, 1964

By W. J. Millard

IT is well known that the excellent mica from Brazil played a most vital role in World War II. Increased production from Brazil was necessary and with the assent of the Brazilian Government engineers from the United States went down to study the problem. One of the means advocated by a former Chief of Metals and Minerals Branch of the Board of Economic Warfare, Mr. Carl Schwegler, was the use of stripping machinery. On a pegmatite dike favorable for stripping, a small number of miners could produce as much mica per day as an underground mine employing many times the same number of men. TYPE OF DEPOSIT FAVORABLE The most favorable type of pegmatite for stripping is that which lies more or less parallel to the surface of the slope of a hill. Pegmatite dikes that cut directly into a hill are not suitable for stripping except where a long outcrop extends along the slope. Then just the outcrop could be stripped to some advantage. The ratio at no time should be greater than I:3; i.e., I cu. m. of pegmatite to 3 cu. m. of waste, unless the pegmatite was extremely rich throughout the dike. THE ECONOMICS OF STRIPPING When a mineral is badly needed to win a war, not too much heed can be paid to costs. The object is to produce rapidly and in quantity. Again, a deposit containing a superior grade of mica may be worked although the pegmatite is more costly to work than another deposit containing a considerably greater amount of poor-grade mica. For example: one deposit may have a little mica that will run II per cent good stain and better when trimmed and contain 2 kg. of crude mica per cubic meter of pegmatite, while another may run 0.2 per cent of good stain and better, and have a mica content of 10 kg. per cubic meter. A number of stripping projects were found in the vicinity of Raul Soares, Minas Gerais. These were sampled in the various old workings by cutting out a cubic meter of pegmatite and saving the usable mica. Because of the proximity to the surface of most stripping projects, about 34 per cent only of the mica is good for trimming. The other 66 per cent is too badly weathered to be of use. In Table I are given data collected on five properties near Raul Soares. In Table I, "value per cubic meter of pegmatite" means the value of the qualified mica obtainable. This does not include the half trimmed and splittings. By "value per cubic meter of matter" is meant the value of the qualified mica per cubic meter of material excavated; i.e. both waste and pegmatite. This is the figure that should he used to determine whether the operation will be profitable or not, as the cost of stripping is about $0.I4 per cubic meter. With new machinery not subject to breakdowns, the cost per cubic meter should be around $0.10 to $0.I5. To this should be

Jan 1, 1946

By W. J. Millard

It is well known that the excellent mica from Brazil played a most vital role in World War 11. Increased production from Brazil was necessary and with the assent of the Brazilian Government engineers from the united' States went down to study the problem. One of the means advocated by a former Chief of Metals and Minerals Branch of the Board of Economic Warfare, Mr. Carl Schwegler, was the use of stripping machinery. On a pegmatite dike favorable for stripping, a small number of miners could produce as much mica per day as an underground mine employing many times the same number of men. Type of Deposit Favorable The most favorable type of pegmatite for stripping is that which lies more or less parallel to the surface of the slope of a hill. Pegmatite dikes that cut directly into a hill are not suitable for stripping except where a long outcrop extends along the slope. Then just the outcrop could be stripped to some advantage. The ratio at no time should be greater than 1: 3; i.e., I cu. m. of pegmatite to 3 cu. m. of waste, unless the pegmatite was extremely rich throughout the dike. The Economics of Stripping When a mineral is badly needed to win a war, not too much heed can be paid to costs. The object is to produce rapidly and in quantity. Again, a deposit containing a superior grade of mica may be worked although the pegmatite is more costly to work than another deposit containing a considerably greater amount of poor-grade mica. For example: one deposit may have a little mica that will run 11 per cent good stain and better when trimmed and contain 2 kg, of crude mica per cubic meter of pegmatite, while another may run 0.2 per cent of good stain and better, and have a mica content of 10 kg. per cubic meter. A number of stripping projects were found in the vicinity of Raul Soares, Minas Gerais. These were sampled in the various old workings by cutting out a cubic meter of pegmatite and saving the usable mica. Because of the proximity to the surface of most stripping projects, about 34 per cent only of the mica is good for trimming. The other 66 per cent is too badly weathered to be of use. In Table I are given data collected on five properties near Raul Soares. In Table I, "value per cubic meter of pegmatite" means the value of the qualified mica obtainable. This does not include the half trimmed and splittings. By "value per cubic meter of matter" is meant the value of the qualified mica per cubic meter of material excavated; i.e. both waste and pegmatite. This is the figure that should be used to determine whether the operation will be profitable or not, as the cost of stripping is about $0.14 per cubic meter. With new machinery not subject to breakdowns, the cost per cubic meter should be around $0.10 to $0.15. To this should be

Jan 1, 1948

By W. J. Millard

It is well known that the excellent mica from Brazil played a most vital role in World War 11. Increased production from Brazil was necessary and with the assent of the Brazilian Government engineers from the united' States went down to study the problem. One of the means advocated by a former Chief of Metals and Minerals Branch of the Board of Economic Warfare, Mr. Carl Schwegler, was the use of stripping machinery. On a pegmatite dike favorable for stripping, a small number of miners could produce as much mica per day as an underground mine employing many times the same number of men. Type of Deposit Favorable The most favorable type of pegmatite for stripping is that which lies more or less parallel to the surface of the slope of a hill. Pegmatite dikes that cut directly into a hill are not suitable for stripping except where a long outcrop extends along the slope. Then just the outcrop could be stripped to some advantage. The ratio at no time should be greater than 1: 3; i.e., I cu. m. of pegmatite to 3 cu. m. of waste, unless the pegmatite was extremely rich throughout the dike. The Economics of Stripping When a mineral is badly needed to win a war, not too much heed can be paid to costs. The object is to produce rapidly and in quantity. Again, a deposit containing a superior grade of mica may be worked although the pegmatite is more costly to work than another deposit containing a considerably greater amount of poor-grade mica. For example: one deposit may have a little mica that will run 11 per cent good stain and better when trimmed and contain 2 kg, of crude mica per cubic meter of pegmatite, while another may run 0.2 per cent of good stain and better, and have a mica content of 10 kg. per cubic meter. A number of stripping projects were found in the vicinity of Raul Soares, Minas Gerais. These were sampled in the various old workings by cutting out a cubic meter of pegmatite and saving the usable mica. Because of the proximity to the surface of most stripping projects, about 34 per cent only of the mica is good for trimming. The other 66 per cent is too badly weathered to be of use. In Table I are given data collected on five properties near Raul Soares. In Table I, "value per cubic meter of pegmatite" means the value of the qualified mica obtainable. This does not include the half trimmed and splittings. By "value per cubic meter of matter" is meant the value of the qualified mica per cubic meter of material excavated; i.e. both waste and pegmatite. This is the figure that should be used to determine whether the operation will be profitable or not, as the cost of stripping is about $0.14 per cubic meter. With new machinery not subject to breakdowns, the cost per cubic meter should be around $0.10 to $0.15. To this should be

Jan 1, 1948

By AIME AIME

ONE hundred and eighty-three A.I.M.E. members, W. S. Morris with an excellent technical program, John Suman, "oil" President of the Institute, and a Texas norther blew into Dallas for the fall meeting of the Petroleum Division at the Adolphus Hotel, Oct. 16-18. and the most successful result will give future meeting committees a high mark to equal. People were still talking of the terrific heat the week before and it was quite noticeable that with the lowered temperatures, the members were in fine fettle for the technical sessions. The total attendance on Friday noon was

Jan 1, 1941

By F. Louis Grammer

The difference between blast-furnace gas and ordinary producer-gas is chiefly that blast-furnace ga,s is higher in CO2 and lower in hydrocprbons and hydrogen, as is shown in Table I. Table I.— Volume-Composition of Producer-Gas and Blast-Furnace Gas. Producer-Gas. Blast-Furnace Gas. Per Cent. Per Cent. (Volume.) (Volume.) CO......22 to 33 25.5 CO2........ 6 to 3 13.0 H...13 to 7 1.5 CH4........ 3 to 1.5 0.0 N..54 to 60 60.0 If natural gas, water-gas, or by-product coke-oven gas be available, blast-furnace gas can be made to resemble very closely in calorific value the average producer-gas by merely mixing in one of these products, which are richer in hydrocarbons and hydrogen. Blast-furnace gas, however, possesses a virtue which is sometimes absent in producer-gas, especially with coals very high in sulphur—namely, that blast-furnace gas is relatively free from sulphur. In some regions the sulphur in coal is an acute problem, and it is highly desirable to find uses for the mine-products which are specially high in sulphur-content. An advantage would be gained if the coal were coked before charging into the gas-producers, but this adds to the expense. Sulphur may be partly eliminated from coal by washing it, the success depending upon the form in which the sulphur occurs, and upon the efficiency of the washer, but it cannot always be removed to the extent desired at most American plants. It is not practicable to remove all the sulphur from producer-gas by washing it, though Mr. Reinhardt, of Germany,

Jan 1, 1909

By H. B. Aaron

DIFFRACTION micrography provides a useful tool for studying complex strain fields. Newkirk1 observed an X-ray diffraction effect due to strain interactions at the intersection of slip lines in LiF and attributed this effect to a reduction of primary extinction due to strain-gradients. It is the object of the current study to correlate the observations and rationalize the appearance of the strain fields around intersecting slip planes in x-ray extinction contrast studies by a semiempirical method. Fig. 1 shows typical strain patterns observed by Newkirk on (200) reflections in LiF using chromium radiation and a variation of the Berg-Barrett (B-B) technique. In other reflections (e.g., {220}) the strain patterns are absent. If one defines a vector 3 along the striations from the more to the less dense portion, and a vector v in the direction of the X-ray beam parallel to the reflecting plane, such that both 3 and v lie in the reflecting plane of the crystal, Fig. 1, then the positions of the striations on (200) reflections appear to be confined such that where is the angle between v and 3, Fig. 1. It will be seen later that 3 is the projection of Sb on the reflecting planes. In some instances the strain effects are observed only above the slip plane, Fig. l(C), in others only below, Fig. l(D). Where the strains appear to be both above and below the slip line, Fig. l(E), one sees that there are actually two slip lines present, Fig. 1(F) and (G). Furthermore, the strain patterns are observed in the vicinity of the intersection of emergent edge dislocations (diagonal lines on the micrograph) with emergent screw dislocations (horizontal lines on the micrograph), e.g., Fig. l(HJ) and (MN), but not in the vicinity of the intersection of two emergent edge dislocation arrays, Fig. l(HJ) and (KL). However, in the latter case the contrast of the dislocation array itself is reduced along (KL) and increased along (HJ). Line (AB) in Fig. 1 is the edge of the crystal, and the material to the left of (AB) is the side of the sample. Any empirical correlation will have to include most, if not all of the aforementioned observations. Lang2 and Newkirk1 determined that the condition for contrast is given by g-b#0 [2] where b is the Burgers' vector and g the diffraction vector (normal to the reflecting planes). If one assumes that the direction of the observed strain fields is given by the vector sum of the slip vectors (Sb1) of the intersecting slip planes, then Eq. [2] may be generalized to g-S#0 [3]

Jan 1, 1963

By E. D. Levine

A study was made to delineate the dislocation mechanistns controlling prismatic and basal slip at low temperatu.ves in titanium containing approximately 100 ppm interstitial impurities. Mechanical tests were performed on single crystals prepared from electron-beam zone-refined crystal bar by a thermal cycling technique. Measurements were made of critical shear stress and activation volume as functions of temperature. The study revealed that, for prismatic slip, two overlapping thermally activated processes aye rate-controlling below 300°K. From absolute zero to 220°K, reasonable agreement is obtained between experiment and the mechanism which involves overcoming 01. the Peierl's stress by the formation of double kinks in dislocation lines lying in close-packed directions. Between 220°and 300°K prismatic slip is controlled by a difierent thermally activated process which is characterized by a stronger temperature dependence of critical shear stress. Above 300OK, prismatic slip is controlled by an athermal process. Basal slip is controlled by a thermally activated process between 195' and 400°K. An approximate analysis suggests that it is the Peierl's mechanism that is rate-controlling in this range. Below 195°K, the occurrence of mechanical twinning obscures the features of basal slip. Above '400 X, basal slip is controlled by an athermal process. TITANIUM, an hcp metal with a less-than-ideal axial ratio of 1.587, deforms primarily by {1010}( 1210) slip at low temperatures.'-3 The critical shear stress is strongly affected by the presence of interstitial impurities. For example, churchman3 reported that a decrease in the combined nitrogen plus oxygen content from 0.1 to 0.01 pct decreased the critical shear stress for prismatic slip from 9.19 to 1.4 kg per sq mm at room temperature. Basal slip, however, is less strongly affected by interstitials; Churchman's data for {0001}( 1210) slip show a smaller decrease in critical shear stress from 10.90 to 6.3 kg per sq mm at room temperature, for the same decrease in interstitial content. In the present study, an attempt has been made to delineate the dislocation mechanisms controlling prismatic and basal slip in titanium. In view of the strong effects that impurities may have on at least some of these mechanisms, experiments were performed on material of reasonably high purity, so that comparisons could be made with investigations on material of both lower and higher impurity contents. EXPERIMENTAL PROCEDURE Single crystals were prepared from 1/4-in.-diam electron-beam zone-refined crystal bar, furnished by Materials Research Corp. The analysis of the zone-refined material is given in Table I, indicating that the interstitial content is of the same order as the purer of the two materials investigated by Churchman. The Vickers diamond pyramid hardness number of this material was 85.6. This value agrees quite well with the correlation between hardness and interstitial content proposed by Ogden and Jaffee.4 On the basis of hardness, the material investigated in this study is, therefore, intermediate in purity between that studied by Churchman (Vpn = 97) and that studied by Spangler and Herman (Vpn = 62.4).5 Single crystals were grown by thermal cycling through the a-ß transformation. Two-in.-long by 1/2-in.-diam bars were heat-treated in vacuum for 4 hr at 1160°C. The temperature of the furnace was then lowered to 865°C, and the specimens were maintained at this temperature for 7 days. This procedure was repeated several times, resulting in the growth of individual crystals up to 14 in. in length. Specimens for testing in direct shear along either {l010}(1210) or {0001}(1210) were obtained by electrical-discharge machining. The specimens were in the form of cylinders 1/4 in. in diam by approximately 1/2 in. in length, with a reduced section 1/8 in. in diam and a in. in length. All specimens were electropol-ished subsequent to electrical-discharge machining. The ends of the test specimens were cemented in stainless-steel collars. Shear tests were conducted in an Instron Tensile Tester at various temperatures from 4.2° to 508°K. Measurements were made of critical shear stress and activation volume. The latter measurements were performed by measuring the change in flow stress upon cycling the strain rate between ?1 = 2.62 x 10-3 per sec and ?2 = 2.62 x 10-4 per sec. Activation volume, I)*, is defined as u* =ßkT [11

Jan 1, 1967

By C. H. Hewitt, J. T. Morgan

The Fry cocurrent in situ combustion project was carried out in a 3.3-acre portion of a lenticular body of Robinson sandstone of Lower Pennsylvanian age. This particular sand body is about 12,000 ft long and 3,500 ft wide; it varies in thickness from 0 to 55 ft, trends northeast-southwest, and occurs at depths between 880 and 936 ft. The reservoir is made lip of three distinct types of sandstone, each having characteristic reservoir properties. (See Table I.) Textural gradients, sedimentary structures, erosional basal contact, absence of marine fossils, and geometry of the sand body all indicate that the sandstone is of fluvial origin and part of an extensive complex of river-deposited sedirnents. Geographically oriented cores reveal that the direction of depositing currents was to the southwest, parallel with the sand-body trend. Directions of maximum pernleability also parallel the northeast-southwest trend. The layered nature of the sandstone reservoir and its directional properties have controlled the areal distribution of produced gas, the ignition behavior, and the vertical profile and horizontal extent of the conbustion zone. INTRODUCTION This paper is the first of a series of three papers that describe Marathon Oil Co.'s Fry in situ combustion test. This first paper contains the results of a detailed study of the geology and reservoir characteristics of the Robinson sandstone reservoir in which the Fry test was conducted. The second and third parts of this series report on the field operations' and performance. The geologic work reiated to cores cut through the combustion zone, and to the effect of reservoir properties on the combustion process is included in the third paper.' GEOLOGIC STUDY OF THE FRY RESERVOIR The close spacing of continuous cores in the Fry combustion site provided an unusual opportunity to study the details of distribution of lithology and reservoir properties within a petroleum reservoir. The objectives of the study were: 1. To determine the distribution of rock types within the Robinson sandstone in the Fry area; 2. To determine the relationship of porosity and permeability to rock properties and establish the distribution and homogeneity of reservoir properties; 3. To determine the origin of the Robinson reservoirs so that attempts could be made to predict the size, shape, distribution, extent and continuity of Robinson reservoirs; and 4. To determine what effects, if any, the reservoir rocks might have on the combustion process, so that the process might be better evaluated. AVAILABLE MATERIALS Continuous cores were cut from 14 wells in the vicinity of the Fry site. Ten of the cores are from within the 3.3-acre pattern; six of the 10 were cut prior to combustion, and four were cut through the combustion zone. Cores from the Emma Fry No. 15, Wampler 0-2 and Fry B wells (Figs. 1, 2 and 9) were geographically oriented by the Eastman Oil Well Survey Co. Maps, electric

Jan 1, 1966

Graham Macfarlane, Louisville, Ky. (communication to the Secretary*):—In the paper by S. Whinery under the above caption mention is made of the occurrence of Clinton iron-ore in Kentucky and Tennessee. I believe that this is really a ferruginous limestone in what the Tennessee Geological Survey calls the Fernvale formation. In Davidson and Sumner counties, Tennessee, this occurs at about the horizon stated by Mr. Whinery. The maximum thickness of the deposit is from 10 to 12 ft., but it averages between 6 and 8 ft., lying practically horizontal. The Red River Furnace Co., under my direction, opened a mine in the Fernvale " red rock" near Goodlettsville, Tenn., several years ago. When it was first found we assumed it to be Clinton ore, but the absence of the characteristic Clinton fossiliferous structure and other evidence satisfied us that it is really limestone saturated with iron oxide. I think the very low ratio of SiO2 and Al2o3 to iron content as compared with the usual ratio in the Clinton ore is also evidence in favor of its being the Fernvale. The Tennessee Geological Survey states that at outcrops where this ore has been leached it carries about 44 per cent. of iron; under cover, from 20 to 25 per cent. of iron and from 3 to 6 per cent. of silica, the remainder being chiefly calcium carbonate. Its phosphorus content is irregular and generally rather high for a foundry-iron mixture. Since it is more than self-fluxing, we have at times operated with a mixture of brown hematite and red ore with no limestone.

Jan 1, 1913

By E. T. Turkdogan, L. S. Darken, R. J. Fruehan

Further consideration is given to the application of the quadratic formalism to evaluate the thermodynamics of binary metallic solutions from experimental data. The use of the thermodynamic relation, in(r1/r2) dN2= 0, in the evaluation of activity data is discussed. The empirical correlations derived previously for enthalpies and entropies of solutions in Part II are shown to be predicted theoretically for a special case. The applicability of the quadratic formalism to the activity coefficients for binary organic solutions is demonstrated. DARKEN&apos; introduced a quadratic formalism describing the composition dependence of the activity coefficient in the terminal regions of binary metallic solutions. In a subsequent paper (Part II) Turkdogan and Darkeen2 extended this formalism to the enthalpy and excess entropy of binary metallic solutions, and showed that the available data for the activity coefficients and heats of mixing of many solid and liquid binary metallic solutions confirm reasonable applicability of the quadratic formalism. Much variation was noted, however, in the extent of the terminal regions from one system to another, commonly from about 10 to 70 at. pct. For the present purpose, equations derived previously need be given only for one of the terminal regions, e.g. region (I), 0 < N2 < N&apos;2: where subscripts 1 and 2 indicate quantities for components 1 and 2 and superscript ° the values at infinite dilution; other terms are: N = atom fraction, y = activity coefficient, = partial molar enthalpy of solution, ?HM = molar heat of solution, ?Sx = excess partial molar entropy, ?Sx = excess molar entropy. The heat and entropy like terms (constants) el, and S12 are defined in the following expression for the temperature dependence of a12, thus a12 £12 — s12 rnl 2.303RT 2.303 R l J Using this formalism, the thermodynamics of binary systems in the terminal regions may be described over a temperature range of experisental interest in terms of eight quantities: Lº1—Lº2, ?Sx1,?Sxº2 , £12 ,£21,fiai, S12, and S21. In the previous paper, these quantities were evaluated for many solid and liquid binary solutions, and two significant correlations were observed between these thermodynamic quantities pertaining to the terminal regions, thus [10] and [11] Combining Eqs. [9], [l0], and [ll] and rearranging gives a12 -a21 = 1.382R log rº2- 0.08(?S2xº- ?S1xº) [12] Since the second term involving entropies is small and Eqs. 1101 and [ll] are empirical relations, the above equation may be simplified to a12 = 2.74 log rº2 [13] Using the values of a&apos;s and y°&apos;s compiled in Parts I and 11, a12 - a21 is plotted against log(rº2/rº1) in Fig. 1; the line drawn has a slope of 2.74. For Hg-K at 300°C, log(yº2 /yº1)= —5.10, not shown in Fig. 1, and from correlation [13], a12 - a21 = -13.97; this is to be compared with -10.76 derived from the activity data (Part II).

Jan 1, 1970

By Luther Wagoner

Having given in a former paper1 the results of assays of sea-water, bay-mud, dredgings from San Francisco bay, etc., and believing it might be interesting to extend the work to include some deep-sea dredgings, I procured from the Smithsonian Institution six samples taken by the U. S. steamer Albatross, and marked as follows: Depth in No. station. Locality. Latitude. Longitude. Fathoms. 1. 2103. Off Delaware bay. 38&apos; 47&apos; 20" N., 72o 37&apos; W., 1,091. 2. 2265 to Between Chesapeake 35° 37&apos; N., 74O to 75O W., 49 to 70. 2297. bay and Hatteras. 3. 2420. Off Chesapeake bay. 37&apos; 03&apos; 20" N., 74O 31&apos; 40" W., 104. 4. 2528. East of Georges 41° 47&apos; N., 65o 37&apos; 30" W., 677. bank. 5. 2572. SE. of Georges bank. 40° 29&apos; N., 66O 041 W., 1,769. 6. 2681 South of Nantucket. 39&apos; 43&apos; N., 70&apos; 29&apos; W., 990. The above samples were assayed by the cyanide-method, described in my former paper (p. 807), about 30 g. of sample being used, and the following results were obtained: Value in Milligrams Per Metric Ton. No. Station. Gold. Silver. 1. 2103. 145 1,014 2. 2265 to 2297. 44 304 3. 2420. 15 353 4. 2628. 267 1,963 5. 2572. 125 377 6. 2681. 66 414 No. 4, Station 2528, consisting of red clay and volcanic ash, mas re-assayed ; 35 g. was roasted at a low heat for one hour and weighed 34.05 g. after roasting; the color was changed to a deep brick-red; and when assayed by the cyanide-method

Jan 1, 1908