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By Robert E. Green, Robert A. Swanson

Ultrasonic attenuation and stress were measured simultaneously as a function of strain for aluminum single crystals tested in compression. The propagation mode and polarization of the ultrasonic waves were chosen with reference to the theoretical analysis presented in Part I. Good experimental agreement was found with this theory and with earlier experimental work in tensile tests. Support is given to the contention that ultrasonic -attenuation measurements are much more sensitive to the early stages of plastic deformation than are conventional stress -strain measurements. Dislocation meclzanisms are suggested which are compatible with both the present and the earlier experimental observations. A brief historical introduction of previous work concerned with the study of dislocation motion by ultrasonic-attenuation measurements has been given in Part I. The purpose of the present experimental investigation was to make use of the theoretical considerations from Part I in order to gain further knowledge concerning dislocation motions during plastic deformation. Emphasis was placed on the choice of propagation modes and polarization of the waves, such that interactions between these waves and dislocation movement on a specific slip system could be investigated. EXPERIMENTAL PROCEDURE Since the primary concern of this investigation was dislocation damping, other sources of damping were eliminated as much as possible. The test specimens were 0.5-in.-diam single crystals of 99.99+ pct pure aluminum grown from the melt by a modified Bridgman technique. Ultrasonic-attenuation measurements were made at a frequency of 10 megacycles per sec using a Style 56A001 Ultrasonic Attenuation Comparator manufactured by Sperry Products Inc. after a design by Chick, An- derson, and Truell.1 At this frequency the particle vibrations are essentially adiabatic and thermo-elastic damping should be absent. This frequency was selected since it was one of the frequencies used by Hikata et al.2 n conjunction with tensile tests on aluminum crystals, thus allowing direct comparison with this work. The use of quartz-crystal transducers insures that the amplitude of vibration is not sufficient to cause the break away of the dislocation loops from weak pinning points. The quartz-crystal transducers used all possessed a resonance frequency of 10 megacycles per sec and were thin circular discs 0.375 in. in diameter. X-cut crystals were used to generate the longitudinal waves and AC-cut crystals to generate the transverse waves. Aluminum was chosen as the basic material because it is a typical fcc metal whose deformation behavior is well-known and because the most recent work at the time the present research was initiated was that by Hikata et al. who used aluminum. The present tests were run in compression to see if such tests could be satisfactorily run in compression without being severely masked by grip effects.. A second reason was to compare the results in compression with those of the tension experiments of Hikata et at. Finally, shorter specimens could be run in compression with less over-all damping than in tension. All of the aluminum-crystal test specimens used were oriented for plastic deformation by single slip; i.e., they possessed the orientation having a maximum Schmid factor of 0.5. Because of the well-known time dependence 3 of plastic strain at constant stress, the test specimens were subjected to a quasi-static loading. Previous investigations4-' have indicated that attenuation after plastic deformation is also time-dependent, so any continuous loading of the specimen would show that changes in attenuation were also a function of strain rate. By loading the specimen to a given stress level and holding it constant until the recovery process was essentially complete, the time dependence was eliminated. The quartz transducer was permanently bonded to the aluminum crystal using Eastman 910 cement. Excess cement was removed from around the quartz with DMF (dimethyl formamide). Eastman 910 was found to be far superior for bonding, especially with regard to the shear-wave transducer, since it formed a solid bond which readily supports shear-wave propagation. The solid bond formed by the Eastman 910 cement also insured that the quartz

Jan 1, 1964

By P. A. Beck, Hsun Hu

As described by means of quantitative pole figures, the annealing texture of highly rolled aluminum consists of the four retained components of the rolling texture near (123) [121], rather more sharply developed, and of a cube texture component. Local reorientation corresponds fairly well to 40 rotation around a [111] axis. In copper strip rolled 96 pct, the annealing texture is mainly the cube texture, with the four twin orientations as minor components. The annealing texture of highly rolled brass strip consists of four components of the (225) [734] type. REVIEWS by Wassermann,' Barrett,' Dunn,3 and Richards' give detailed accounts of the numerous publications dealing with annealing textures in rolled polycrystalline metals. Although a considerable amount of information of an empirical nature has been available for many years, even relatively recently the state of recrystallization-texture theory was considered uncertain. The reviews referred to above deal mainly with the description of experimental results, without developing generally applicable principles, suitable for explaining the observed annealing textures. The investigations of Burgers and Louwerse5 and Barrett6 showed that in deformed single crystals of aluminum the orientation of the recrystallized grains deviates drastically from that of the deformed matrix. According to Barrett and to Beck and Hu,7 the relationship between the two orientations can be best described as a rotation of about 40° around a [111] axis. Recent investigations led to similar orientation relationships for various face-centered cubic metals with a simple texture, regardless of whether the annealing process was recrystallization (strain-free grains growing in a highly deformed large crystal of aluminum' or copper: or in deformed cube texture copper") or coarsening (strain-free grains growing at the expense of an essentially strain-free, fine grained, polycrystalline matrix of single orientation texture copper,'" aluminum,' or nickel-iron). Barrett" pointed out the possibility that this orientation relationship may be a result of the orientation dependence of the rate of grain boundary migration. More recently, considerable evidence became available in favor of this view.12,13 In apparent contradiction to the above results, it has been known for a long time that the recrystallization texture of polycrystalline metals is often similar to the deformation texture. This was found, in particular, for the fiber textures obtained by wire drawing, compression, or compression rolling.' Following a suggestion by Barrett,2 Kronberg and Wilson and Beck and Hu7 described a mechanism by which a [Ill] fiber texture, resulting from plastic deformation, may be retained upon recrystallization, even though the orientation of individual recrystallized grains differs from that of the local area of the matrix in which they grow. This mechanism assumes that the reorientation corresponds to a rotation around the [1ll] axis, which is also the fiber axis. The present investigation was undertaken to examine in detail the orientation relationship between deformation texture and recrystallization texture in rolled face-centered cubic metals, using the quantitative methods of texture determination that have recently become available.", '" Quantitative pole figures for the rolling textures of such metals were reported recently." Quantitative pole figures for annealing textures are given in the present paper. It was planned to investigate whether or not the orientation relationship previously found for the recrystallization of deformed single crystals, and for coarsening of single-orientation texture material, is capable of accounting for the observed, often complex, annealing textures in rolled polycrystalline metals. The question whether local reorientation is predominant, even in the early phases of the annealing process, was to be studied with high purity aluminum. by making use of an oxide film and sensitive tint illum-

Jan 1, 1953

By E. S. Wheeler

The conversion plant of the Clirnax Molybdenum Co. is at Langeloth, Washington . County, Pennsylvania, approximately 30 miles west of Pittsburgh. The molybdenite concentrates converted originate in the company's mine and mill at Climax, Co1orado.l They are delivered to the Langeloth plant by rail packed in 175-lb. bags. War production of concentrates at the mill started in 1918 but it was not until about 1920 that any peacetime uses of the metal began to develop. Early conversion of Climax concentrates was mainly into ferromolybden um, produced for the company by a custom smelter. Chronology of Plant Development In 1924 a small hand-rabbled roaster was built at Langeloth, the roasted concentrates being used principally in the production of calcium molybdate by a conventional wet method. The size of the original plant is indicated by the fact that during 1925 about 400,000 lb. of molybdenum was converted. In 1925-1926, a multiple-hearth roaster was erected (12-ft. 0.d. 8-hearth Nichols-Herreshoff); also the first experimental ferromolybdenum was produced at Langeloth. At this time sa.1cium molybdate was produced in the furnace by a process to be described. With continued interest in and increased use of molybdenum, it was necessary to arrange for enlargement of plant facilities. Although only one 12-hearth 16-ft. 0.d. roaster was erected at this time (1929), plans were prepared for a duplicate unit, which was completed in 193 5, giving a potential roasting capacity of some 16,000 lb. Mo per day under the then prevailing roasting practice. Shortly after this second unit had gone into production, increased demand overtaxed the plant, and as a result two 12-hearth 18-ft. 0.d. roasters were placed in operation in 1936 and 1937. This gave the plant a potential roasting capacity of 40,000 to 50,000 lb. Mo per day, and necessitated major changes in auxiliary handling equipment as well as improved facilities for producing ferromolybdenum. Following the introduction (1938) of a new product—molybdic oxide briquets— and because of the constantly growing demand for molybdenum, plant facilities were again enlarged in 1940 and 1941 by the erection of two 16-hearth 18-ft. 0.d. roasters. These two new units, along with a change in roasting practice, brought the potential roasting capacity to 150,ooo to 190,000 Ib. Mo per day, based on about 15 lb. per sq. ft. of hearth area. Other departments of the plant have necessarily grown, until today the plant is modern throughout and completely mechanized, with equipment geared to produce marketable products as follows:

Jan 1, 1944

By E. S. Wheeler

THE conversion plant of the Climax Molybdenum Co. is at Langeloth, Washington County, Pennsylvania, approximately 30 miles west of Pittsburgh. The molybdenite concentrates converted originate in the company's mine and mill at Climax, Colorado.1 They are delivered to the Langeloth plant by rail packed in 175-lb. bags. War production of concentrates at the mill started in 1918 but it was not until about 1920 that any peacetime uses of the metal began to develop. Early conversion of Climax concentrates was mainly into ferromolybden um, produced for the company by a custom smelter. CHRONOLOGY OF PLANT DEVELOPMENT In 1924 a small hand-rabbled roaster was built at Langeloth, the roasted concentrates being used principally in the production of calcium molybdate by a conventional wet method. The size of the original plant is indicated by the fact that during 1925 about 400,000 lb. of molybdenum was converted. In 1925-1926, a multiple-hearth roaster was erected (12-ft. o.d. 8-hearth Nichols-Herreshoff); also the first experimental ferromolybdenum was produced at Langeloth. At this time calcium molybdate was produced in the furnace by a process to be described. With continued interest in and increased use of molybdenum, it was necessary to arrange for enlargement of plant facilities. Although only one 12-hearth 16-ft. o.d. roaster was erected at this time (1929), plans were prepared for a duplicate unit, which was completed in 1935, giving a potential roasting capacity of some 16,000 lb. Mo per day under the then prevailing roasting practice. Shortly after this second unit had gone into production, increased demand overtaxed the plant, and as a result two 12-hearth 18-ft. o.d. roasters were placed in operation in 1936 and 1937. This gave the plant a potential roasting capacity of 40,000 to 50,000 lb. Mo per day, and necessitated major changes in auxiliary handling equipment as well as improved facilities for producing ferromolybdenum. Following the introduction (1938) of a new product-molybdic oxide briquets and because of the constantly growing demand for molybdenum, plant facilities were again enlarged in 1940 and 1941 by the erection of two 16-hearth 18-ft. o.d. roasters. These two new units, along with a change in roasting practice, brought the potential roasting capacity to 150,000 to 190,000 lb. Mo per day, based on about 15 lb. per sq. ft. of hearth area. Other departments of the plant have necessarily grown, until today the plant is modern throughout and completely mechanized, with equipment geared to produce marketable products as follows:

Jan 1, 1944

By Harry Tufft

FOR many years the treatment of mine timbers with preservatives was confined to a few pioneer plants in the United States, and it is only in the past few years that the practice has grown appreciably. In a field study of the mine timber situation in Illinois, Pennsylvania, and further study of the situation in the country as a whole, I found that most mine operators did not keep records on mine timber that would afford the necessary data for determining the possible savings from preservative treatment. Some primary questions the mining engineer should be able to answer before making his decision on preservative treatment. for a particular mine are: (1) Quantity and cost of timber going into permanent workings annually; (2) quantity and cost of annual replacements due to decay; (3) life of untreated timber in different places; (4) desired life of timber. With the necessary facts in hand, he can then con-fidently select a method of treatment and preservative that is best suited to his needs. In former years when timber was cheap and an abundant supply could be had close at hand, as was the case in the majority of mining regions, the mine op-erator did not concern himself much except to see that enough was supplied to keep the workings safe. The cost figure for timber and labor, distributed to cost per ton of ore, was all the bookkeeping necessary. With present costs of timber, labor, and increasing length of haul to the mine, many operators are paying closer attention to timbering costs and studying them in. more detail. Some of the larger and more progressive opera-tions now keep detailed' figures that give information on all phases of their timber problems. However, at the majority of mines the data available do not permit such analysis. Checks as to where the different con-signments are placed, the date of setting, and the period

Jan 6, 1927

By Pierre R. Hines

These notes are based on observations made while on a recent trip through the West, for the purpose of studying the practical operation of the ball-mill. The writer takes this opportunity to express his thanks for courtesies extended at the many plants visited as well as for the valuable data received: While there are several types of ball-mill on the market, particular attention will here be given to the diaphragm type, as the open-trunnion type, especially the conical mill, has been thoroughly discussed in the Transactions. There is a prevailing impression that the ball-mill is a recent development; however, ball-mills were used extensively in Montana and other western states for crushing ores for concentration as far back as 1808. Its present prominence is due in part to its recent successful application by one of the large copper companies. Without any reference to dry grinding, the first successful ball-mill for wet crushing, which is still in operation, was built 10 years ago.l This mill, designed by Erminio Ferraris for crushing Sardinian ores for concentration, is of more than passing interest. It embodies the peripheral discharge with grates, large forged-steel balls, and the principal features of the modern ball-mill. The results approach present-day practice, the chief differences being that the mechanical construction has been improved in the modern types. Crushing Action of Ball-mill The action of the balls and the principles of crushing have been studied by several investigators.2 Their conclusions are confirmed by results obtained by the writer in experimenting with a small machine built at

Jan 1, 1918

THE fall meeting of the Petroleum Division of the American Institute of Mining and Metallurgical Engineers will be held in Fort Worth on Oct. 19 and 20. The Texas Hotel will be headquarters. for this meeting, the technical sessions being held in the ball-room of the hotel. During the evening of the nineteenth a smoker will be held at the Fort Worth Club. In working out the program for this meeting, J. B. Umpleby, vice-chairman for production engineering, considered it especially desirable to offer as complete a discussion as possible of the subject of air-gas lift, its theory and practice, and its relationship to gas-oil ratios. The entire first day will be devoted to sessions on this subject. One session of the second day will be devoted to the use of electricity in oil-field operations and the second to the handling of congealing oils and paraffin. In addition to this program there will be one or more sessions when petroleum engineering research problems will be discussed. This meeting of the Division is primarily devoted to production engineering, as will be seen from the pro-gram. During one of the sessions, however, a brief review with forecasts of the production situation will be presented.

Jan 8, 1927

By Howard Scott

INVAR is the preeminent low-expansion metal by virtue of the fact that it can be prepared with a zero coefficient of expansion at atmospheric temperature. This fact suggests that there is little room for improvement in the expansion properties of the low-expansion nickel steels. The inference is true, however, only so far as the expansion at and near atmospheric temperature is concerned, because the expansivity of invar starts to increase rapidly on heating above 100° C. and soon reaches the value of copper. Guillaume1 met this situation by increasing the nickel content above that of invar. By that expedient, alloys having low expansivity under higher temperatures were obtained, but not without a large increase in minimum expansivity. The increase in minimum expansivity with increasing nickel content severely limits the technical applications of these alloys at high temperatures. Prospects of improving the expansion characteristics of the nickel steels is offered by alloying. Search for an advantageous alloying addition by analogy with other alloy systems, however, is not productive because the low-expansion characteristic is unique with that system. The rule of mixtures usually gives a fair value of the expansivity of other solid-solution alloys, but in the iron-nickel systems gives a value of 18 X 10-6 per deg. C. whereas the actual value is 1 X 10-6 per deg. C. Apparently suitable additions are not to be deduced from such ingenuous considerations. Recognition of the fact that the temperature range of low expansivity in nickel steels is terminated with loss of magnetism affords a more promising basis for the selection of an advantageous element. One may expect from that fact that additions of nonmagnetic elements to the iron-nickel alloys are detrimental to the low-expansion characteristic. Indeed, carbon, manganese, silicon, copper and chromium are detri-mental to the expansion properties and no nonmagnetic element is reported to be beneficial. From the same viewpoint, the addition of a magnetic metal offers prospects of improvement. As two of the three ferromagnetic elements are already present in the

Jan 1, 1930

By Howard Scott

Invar is the preeminent low-expansion metal by virtue of the fact that it can be prepared with a zero coefficient of expansion at atmospheric temperature. This fact suggests that there is little room for improvement in the expansion properties of the low-expansion nickel steels. The inference is true, however, only so far as the expansion at and near atmospheric temperature is concerned, because the expansivity of invar starts to increase rapidly on heating above 100" C. and soon reaches the value of copper. Guillaumel met this situation by increasing the nickel content above that of invar. By that expedient, alloys having low expansivity under higher temperatures were obtained, but not without a large increase in minimum expansivity. The increase in minimum expansivity with increasing nickel content severely limits the technical applications of these alloys at high temperatures. Prospects of improving the expansion characteristics of the nickel steels is offered by alloying. Search for an advantageous alloying addition by analogy with other alloy systems, however, is not productive because the low-expansion characteristic is unique with that system. The rule of mixtures usually gives a fair value of the expansivity of other solid-solution alloys, but in the iron-nickel systems gives a value of 18 X 10-6 per deg. C. whereas the actual value is 1 X 10-6 per deg. C. Apparently suitable additions are not to be deduced from such ingenuous considerations. Recognition of the fact that the temperature range of low expansivity in nickel steels is terminated with loss of magnetism affords a more promising basis for the selection of an advantageous element. One may expect from that fact that additions of nonmagnetic elements to the iron-nickel alloys are detrimental to the low-expansion characteristic. Indeed, carbon, manganese, silicon, copper and chromium are detrimental to the expansion properties and no nonmagnetic element is reported to be beneficial. From the same viewpoint, the addition of a magnetic metal offers prospects of improvement. As two of the three ferromagnetic elements are already present in the

Jan 1, 1930

By Joseph T. Singewald

AMONG the genetically interesting iron ores of the United States are those of the St. Francis Mountains near Ironton and Iron Mountain, Missouri. They are specular hematite in porphyry. The Iron Mountain and Shepherd Mountain deposits are generally spoken of as veins and those of Pilot Knob, Cedar Hill, and Russell Mountain as bedded deposits. Iron Mountain has produced nearly 4,000,000 tons of ore and Pilot Knob about 1,500,000 tons. The output of the other three localities has been insignificant. Pilot Knob is no longer worked as an iron mine, but the rock is now being crushed for use as road metal. Iron Mountain has continued as an active producer. In 1927, the property was acquired by the M. A. Hanna Co. and an extensive prospecting and exploration campaign was undertaken, so that the ultimate production of Iron Mountain will doubtless greatly exceed the figures mentioned. Previous Theories of Origin Descriptions of the deposits and theories of their origin have been concerned mainly with Iron Mountain and Pilot Knob, which represent the two contrasting types of occurrence. The Iron Mountain deposits have been considered of close magmatic origin by Whitney' in 18.54, Geijer2 in 1915, and Spurr3 in 1927. Whitney referred to them as of eruptive origin. Geijer considered them fissure fillings at magmatic temperatures. Spurr called them vein dikes formed from ore magma which was essentially fluid iron oxide. Schmidt4 in 1872 ascribed them to fissure filling by thermal chalybeate waters of meteoric origin. Nason5 in 1892 advanced essentially the same explanation. One of the best

Jan 1, 1929

By G. Kinder, R. Gilges, W. Stoker, R. Davey

The Bingham Canyon Mine, located 48 km (30 mi) southwest of Salt Lake City, Utah, is operated by Kennecott Copper Corporation's Utah Copper Division. This operation is the world's largest open-pit copper mine and produces 96 000 tons (106,000 st) ore and 327 000 tons (360,000 st) waste each day. This material is loaded by large electric shovels and transported by both truck and rail haulage. To meet the requirements of the Clean Air Act of 1970 (as amended), new facilities for the control of smelter emissions were required and are now nearing the final stages of construction. The facility will be capable of treating all the output from the mine, some 2,530 tons (2,789 st) per day of combined concentrate and precipitate. Costing $280 million and necessitating the replacement of reverberatory furnaces with continuous smelting vessels, the plant, with its massive and complex gas handling and acid-making facility, is required to remain on-stream with the minimum of pollution control equipment outages to conform with Federal Regulation 52.2325.

Jan 8, 1978

By Stanley D. Michaelson

The selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. Development of Technique Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.' Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a 12 by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1947

By Stanley D. Michaelson

The selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. Development of Technique Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.' Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a 12 by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1947

By Stanley D. Michaelson

THE selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. DEVELOPMENT OF TECHNIQUE Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.1 Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a I a by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made, to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1945

By B. C. Craft

Development and exploratory work in Mississippi during 1934 was rather active, resulting in the expansion of the proven area and the drilling of a number of important wildcat wells. Mississippi showed a decrease in drilling operations during 1934. There were 36 completions from 18 counties during this period compared with 58 completions from 22 counties during 1933. There was a considerable decrease in activity in the Jackson gas field in Hinds and Rankin counties with only 9 completions, resulting in 8 gas wells and 1 dry hole compared with 17 gas wells, 2 oil wells and 7 dry holes during 1933. Leasing and geophysical work was reported from several sections of Mississippi last year. Extensive exploration and drilling were carried on in Lamar, Hancock, Pearl River, Greene and Stone counties and on the flanks of the Jackson anticline. Exploration will be continued in these areas during 1935. Four important wildcats were completed during 1934, all of which were reported as being structurally high, but without having important shows of oil or gas. These tests were: Hix and Germany's, J. J. Newman Lumber Co. No. 1, sec. 21-4N.-15W., Lamar County, abandoned at 3520 it. in the Wilcox; United Gas Public Service Company's Davis No. 1, sec. 30-1N.-5W., Greene County, abandoned at 7439 ft.; Gulf Refining Company's May L. Williams No. 1, sec. 33-5S.-17W., Pearl River County, abandoned at 4800 ft.; and Clarke County Oil Company's S. B. Kirkland No. 1, sec. 10-2 N.-16E., abandoned at 4002 ft. in the Eutaw. Another interesting wildcat was Gulf Refining Company's Dantzler Lumber Co. No. 6, sec. 22-4S.-llW., Stone County, abandoned at 7700 ft. (according to reports below the Selma chalk). A total of 114 of the 132 gas wells completed in the Jackson gas field are still producing and have a combined open-flow capacity of 4,145,-696,000 cu. ft. Four of the abandoned gas producers were never connected to the lines. The open-flow capacity of three wells in the Amory field is approximately 1,000,000 cu. ft. Producing against a line pressure of 50 lb., the demand during December greatly exceeded the capacity.

Jan 1, 1935

By Jei Y. Choi, P. G. Shewmon

Using a scratch smoothing technique, the magnitude and the ratio of the surface diffusion coefficients (D,) in two orthogonal directions have been measured on copper for over 100 different surface orientations. Anisotropy was found for all surface orientations except those lying near the (100) and (111); however, these anisotropies were rarely greater than 2.0. The direction of maximum D, appears to coincide with the direction of step edges on any given surface, and if no such steps exist, no anisotropy is found. Thus the jump frequency along a step edge is shown to be faster than the frequency of a jump over the step. The values of Ds and AH for the (100) and the (111) orientations agree quite well with those obtained earlier by the grain boundary grooving technique. It is proven that the directions of maximum and minimum D, (the principal axes) for any surface orientation are known if the perfectly ordered plane corresponding to the given surface orientation possesses a plane of mirror symmetry. Thus the principal axes of D, for orientations lying along the sides of the unit stereo-graphic triangle are always known independent of the extent of thermal disorder or impurity adsorption. The merits and shortcomings of King and Mullins' recently published equations relating Ds to the rate of scratch smoothing are discussed. 1N an earlier paper the authors have reported the variation of the surface diffusion coefficient (Ds) with surface orientation on copper.' In that study Ds was determined from the rate of growth of grain boundary grooves on symmetric bicrystals. This gave values of Ds from 850° to 1070°C in a given direction on each of the six surfaces. It was found that the activation energy for surface diffusion was essentially the same for all six surfaces. The theory developed to explain these data assumed that the free energy of any given type of surface site depends only on the number of nearest neighbors the atom has. This same theory predicted that on any given surface Ds should be independent of direction. The primary aim of this study was to check the validity of this prediction for a wide range of surface orientations on copper. The growth of grain boundary grooves is ideal for studying Ds in a given direction on a given surface over a range of temperatures, because the same specimen can be repolished and used again. However, there is considerable work involved in growing a symmetric bicrystal, and for a survey of Ds in two or more directions on many surface orientations, the technique would be arduous. A better technique appeared to be the scratch method used primarily by Mykura. 2 If a scratch is placed in a surface whose surface tension is independent of orientation, the surface will smooth out in time under the influence of surface tension. To determine D, from this flattening, Blakely and Mykura2 used an approximate solution which was assumed to describe the rate of flattening. Recently King and Mullins3 (hereinafter K and M) have given a more exact analysis of the flattening of an isolated scratch. A second aim of this paper was to check the validity of this newer analysis and compare the values of Ds determined thereby with the values obtained from grain boundary grooving techniques. EXPERIMENTAL PROCEDURE In an effort to work with a particular predetermined surface orientation, light orthogonal scratches

Jan 1, 1963

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Jan 1, 1961

By S. W. Miller

DURING the examination of welds made in steel by the oxy-acetylene and electric-arc processes, the writer has met with some unusual structures, which he has not encountered elsewhere. They seem to be quite closely associated with the welding methods used, and characteristic of them, and would seem to be worth noting, especially as the references to them in the literature are few, and their explanation not clear or complete. An introductory word or two about fusion welding may not he out of place. The oxy-acetylene process is based on the great and very local heat generated by the burning of acetylene in the presence of commercially pure oxygen, the two gases being thoroughly mixed before combustion. The theoretical temperature is somewhat under 4000° C., but on account of various losses, the actual maximum temperature is probably about 3000° C. 1 The electric-arc process, using a metal electrode, which is also the filling rod, generates a still higher and more local temperature. The actual temperature of the positive crater of the carbon arc is given as 3500° C.2 No figures have been found for the temperature of the arc using a metal electrode, but it would probably not be very different from that of the carbon arc, and probably higher than that of the oxy-acetylene flame. The greater localization of the electric arc is due to the fact that it is not surrounded by a large envelope flame as well as to its higher temperature; while the oxy-acetylene white cone, or welding flame, in which the carbon of the acetylene is burned to CO, is surrounded and preceded by -a long envelope flame, itself of quite high temperature, in which the CO is burned to C02, and the hydrogen of the acetylene to

Jan 2, 1918

By K. G. Kreider, J. H. Brophy, J. Wulff

The technology of nickel-activated sintering of tungsten powder has been successfully applied to the densification of plasma-sprayed tungsten. Nickel was added by infiltration in a zinc solution followed by evaporation of the solvent. After sintering one hour at 1300°C density 95 pct of theoretical and transverse rupture strength of 74,000 psi were obtained. Shrinkage was found to be anisotropic and the mechanism of densification was comparable to that found in the nickel-activated sintering of tungsten powder. 1 HE use of a plasma spray gun for the fabrication of massive tungsten parts has become increasingly interesting. Applications now exist where a deposit in the as-sprayed condition is satisfactory. However, these deposits are generally characterized by a lamellar anisotropic microstructure containing 15 pct porosity of which, typically, two-thirds is open to the surface. Mechanically, the as-sprayed deposits fail at relatively low stress levels with a biscuit-like fracture. As a result of these problems the possibility of improving structure and strength by sintering treatments subsequent to spraying is particularly attractive. Preferably this sintering treatment should be adaptable to large bodies of sprayed metal. The similarity between the as-sprayed tungsten structure and that of a powder compact suggests that the relatively low-temperature activated sintering technique1 might be profitably employed in the densification of plasma-sprayed tungsten. It was the purpose of the present investigation to develop a technique for introducing the nickel-activating agent into the sprayed structure, to evaluate the amount and mechanism of densification obtained as a function of time and temperature, and to obtain an indication of the relative strength before and after sintering. EXPERIMENTAL PROCEDURE Powder used for spraying was purchased from the Wah Chang Corp. in several size fractions ranging from an average size of 4 to 150 . These powders were sized further for an explicit study of the influence of average feed size on densification. All powders were dried at 200°C before use. Spraying was accomplished with a Plasma Flame unit manufactured by Thermal Dynamics Corp. Several modifications of the unit were helpful in conducting the investigation. A variable speed auger feed mechanism coupled with the carrier gas mecha nism facilitated the use of fine particle sizes. A coil of ten turns of copper tubing in series with the arc power and concentrically would around the nozzle improved nozzle life and extended the range of operating currents available. The function of the auxiliary coil was to cause the arc to spin and to prevent impingement at only one point in the nozzle. Normally air sprayed deposits were made with an arc maintained at 400 amp at 50 to 70 v. The arc was blown by a gas mixture containing from 5 pct H, 95 pct N for the finest powder feed sizes ranging to 20 pct H, 80 pct N for the coarsest size. The flow rate was maintained at 100 cu ft per hr NTP through a nozzle of 0.25 in. ID. When apraying in air, the powder stream was directed toward an aluminum substrate for ease of mechanical removal of the deposit. The substrate was cooled by diverting the plasma flame with an air jet, and a second jet was directed on the deposit surface. In this configuration a gun-to-work distance of 2 to 3 in. was found to be satisfactory. Fig. 1 represents a typical as-sprayed deposit micro-structure. Laboratory studies of protective atmosphere spraying were carried out in cylindrical chamber 8 in. in diam by 18 in. in length. In operating the nozzle attached to such a chamber, particular care was required to avoid nozzle burn out due to reduced gas flow. The structure and density of the chamber sprayed deposits varied over wide ranges depending on substrate temperature. For the purposes of this investigation, flat deposits were made approximately 2 in. sq by 3/8 in. thick. From these deposits individual samples were cut an ground to a rectangular shape typically 1 1/2 in. by 1/8 in. sq such that the long dimension was perpendicular to the spraying direction. For the study of shrinkage anisotropy deposits up to one inch thick were produced. From these, rectangular samples were cut having a longer dimension parallel to the spraying axis. Prior to the addition of activating agent, the samples were deoxidized in hydrogen at 800°C for 20 min. No detectable dimensional or microstructural change was observed after this treatment. The addition of nickel was accomplished by infil-

Jan 1, 1963

By E. E. Duncan, G. Derce

REVIEW of the literature on hydrogen in steels indicates that inadequate consideration has been given to the influence of temperature gradients on segregation of this element. Even when segregation has been taken into account, the interpretation of the data has been seriously handicapped by incomplete understanding of the mechanism by which it occurred. Recognition of a mechanism which will be described as "thermal segregation" appears to rationalize many of the observed effects so well that it will be reported here with some preliminary experimental evidence. The complete quantitative evaluation of the factors controlling the mechanism will require extensive experimentation. The operation of the mechanism can be described in relation to fig. 1, which shows the temperature dependence of the solubility of hydrogen in iron at

Jan 1, 1951