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By Jun-Yan Chen, Stefan H. Boshkov

Metal mining in the People's Republic of China has shown great growth during the last three decades. Over 150 new iron ore mines and nonferrous base metal mines have been opened since 1949, althouc.11 most of them of medium and small size. Concurrently, many old mines were enlarged and reconstructed. The production of iron ore has grown from less than 1.5 million tons per year in 1949 to almost 76 million tons in 1978. A corresponding increase in the total annual output in the nonferrous metal mineral industry has also been noted. It is well known that the PRC is very rich in metal mineral resources. The reserves of tungsten, tin, antimony, molybdenum and iron ore are ranked in the top three in the world. Recently, some copper porphyry ore deposits were discovered in Jianxi and Anhui provinces. Some large nickel-copper sulfide deposits were discovered and have been developed in North- West China. Huge iron ore deposits have been found in Liaoning, Hebei and Szechwan provinces. However, the bulk of the iron ores are low grade. The growing metal mineral production has met basic domestic needs since 1949; it has also proven China's big resource potential. Mining methods have also undergone development in the past 30 years. About 80% of the annual output of iron ore comes from open pits, whereas about 58% of nonferrous metal ores come from underground mines. Blasthole open stoping, sublevel caving and induced block caving are widely used in Chinese underground metal mines (table 1). This paper summarizes in detail the planning for and application of these three bulk mining methods in the nonferrous and iron ore mines in China. CLASSIFICATION OF UNDERGROUND MINING METHODS The main principle of the classification of underground mining methods in PRC is ground support in the stoping cycles of the mining blocks or stopes. Table 2 lists this classification. It is essentially the same as that given in the SME Mining Engineering Handbook. Be- cause pillar removal is a secondary mining operation, using a variety of methods, they are not included in this classification. As is the case with most classifications in science and technology, there is no precise division between the classes of mining methods. Because of the rapid development of consolidated filling techniques in the past twenty years, some types of open stoping methods use fill in order to improve the ground conditions for the secondary recovery stage. Thus, these mining methods have features of both methods using no artificial support and those using artificial support. In this classification, this type of mining method is placed in the latter category. The shrinkage stoping method has been employed extensively to mine steeply dipping narrow veins of tungsten and tin in the Jiangxi, Hunan, Guangdong and Guangxi provinces. Shrink- age stoping is placed under the class of mining methods without artificial support be- cause this reflects the nature of the Chinese experience. The function of the broken ore serving to support the footwall and/or hanging wall is limited, because the expectation is for the host rocks to remain unsupported until the stope is drawn empty. APPLICATION OF BULK MINING METHODS Bulk mining methods consist of blasthole open stoping, sublevel caving and block caving. These stoping methods have much in common, such as: bulk ore materials blasted by long holes; higher production rates of stopes or blocks and higher production efficiencies of the whole mining systems.

Jan 1, 1981

By M. Kaufman, A. E. Palty

The phase structure and aging characteristics of two nickel-base alloys, Inconel 718 and 702, were investigated. Wrought and cast Inconel 718 showed ArisCb as the major hardening phase, as well as ture was positively identified. Inconel 702 had Ni3Al as the major hardening phase, as well as Cr7C3-of the phase amounts with temperature was determined. 1 HE value of phase structure and aging rate information in the utilization of alloys has been well established. As part of a continuing program, the results obtained on two nickel-base alloys, hconel 718 and Inconel 702, are presented here. Inconel 718 is a relatively new alloy considered for application in both sheet and cast form. Its novelty lies in the use of high columbium tantalum levels in a Ni-Cr-Fe base. No previous work on this type of alloy has been reported in the literature. Inconel 702 is being used now as an oxidation-resistant sheet alloy. Occasional fabrication problems spurred work on its basic structural behavior. The composition of 702 is essentially nichrome plus aluminum and titanium for strengthening and added oxidation resistance. This puts it in the basic class of other nickel-base alloys which have already been studied. The two alloys will be treated separately. INCONEL 718 Phase Study Methods. The phases present after the various conditions investigated were determined by X-ray diffraction analysis of two different elec-trolytically extracted residues and microscopic examination. The extraction methods, using 10 pct HC1 in alcohol and 10 pct H3PO, in water and the X-ray technique have been described previously.' The effect of temperature on the phase behavior was studied on sheet material using the following heat treatments: 2250°F, 2 hr, ice brine quench 1300 100 hr, water quench 2250°F, 2 hr, ice brine quench 1400°, 100 hr, water quench 2250°F, 2 hr, ice brine quench 1500°F, 100 hr, water quench 2250o 2 hr, ice brine quench 1700°F, 48 hr, water quench 2250°F, 2 hr, ice brine quench 1850°, 24 hr, water quench 2250°, 2 hr, ice brine quench 2000°F, 6 hr, water quench The high solution temperature was an attempt to dissolve as many phases as possible without causing melting. Aging exposures were made longer than normal heat treating times in order to approach equilibrium conditions more closely. Cast material was examined only in the as-cast condition and with the normal heat treatment (1700°F, 1 hr, air cool 1325oF, 16 hr,air cool). Materials. The full phase study was performed on sheet specimens approximately 1/2 in. by 2 in. by 1/16 in. The cast pieces were cut from the gating system of a vacuum-cast test piece. An all weld-metal specimen was made by running many weld beads on the sheet using filler made from the sheared edges of the sheet. All external contamination was removed by belt-sanding followed by an electro-polishing treatment. The typical composition of wrought Inco 718 is as follows: The alloy is air-melted. Cast hco 718 may have higher molybdenum. Due to the expectation of finding a Ni-Cb phase (Ni, Cb) for which no X-ray diffraction pattern was available in the ASTM Index, a special 5 Ib heat to the stoichiometric composition of this intermetallic

Jan 1, 1962

By R. I. Jaffee, F. C. Holden, H. R. Ogden

ECENT papers dealing with the properties of unalloyed iodide titanium have been directed primarily at the determination of base-line properties for alloy investigations. Early work was limited to a few tests because of the limited availability of iodide titanium at the time. In the results of papers by Campbell et al.,1 Gonser and Litton,2 Jaffee and Campbell,3 inlay and Snyder,4 and Jaffee, Ogden, and Maykuth, data on mechanical properties are presented for unalloyed iodide titanium in the annealed and cold-worked conditions. Data are presented in this paper which show the effects of heat treatment on the structure and mechanical properties of commercially produced iodide titanium. Correlation is made between microstruc-tural variables and the mechanical properties. Experimental Procedures Melting Stock: The melting stock used was as-deposited iodide titanium, produced by New Jersey Zinc Co. The furnished analysis showed the following range of impurities: N, 0.004 to 0.008 pct; Mn, 0.005 to 0.013; Fe, 0.0035 to 0.025; Al, 0.013 to 0.015; Mo, 0.0015; Pb, 0.0045 to 0.0065; Cu, 0.0015 to 0.002; Sn, 0.001 to 0.01; Mg, 0.0015 to 0.002; and Ni, 0.003. Hydrogen content as determined by vacuum-fusion analysis was 0.0091 wt pct (0.44 atomic pct) after arc melting and fabrication. Nitrogen analysis on the arc-melted and fabricated titanium showed a content of less than 0.002 pct N. The average hardness of the furnished stock was Rf 70, or approximately 85 VHN. Melting Procedure: The as-deposited rods were rolled, sheared, and degreased in preparation for arc melting. The charge was arc-melted with a tungsten electrode in a water-cooled copper crucible under a positive pressure of high purity (99.96 pct) argon. The final ingot was approximately 2 in. in diameter and showed no increase in hardness over that of the initial stock. Fabrication: Heating for fabrication was done in air. It was begun by forging the ingot into a 3/4 in. diam rod, at an initial temperature of 1600°F. Scale was removed by sandblasting. The rod was then swaged to 1/4 in. diam at room temperature through a series of 20 dies, with approximately 10 pct reduction in area between each die. An anneal of 1 hr at 850 °C in air was given after the 1/2 in. die, such that the final cold reduction was 75 pct. Sections cut from this rod were used for test and microstructure specimens. Heat Treatment: Heat treatments were carried out in resistance tube furnaces with stainless-steel linings, under an atmosphere of gettered argon. As further protection against contamination, the specimens were packed in titanium turnings in a titanium sleeve. Control experiments have shown negligible hardness increases with this method, indicating that contamination from oxygen and nitrogen is slight. Three cooling rates were employed in this work; these have been designated as water quenching, argon cooling (to simulate air cooling under a controlled atmosphere), and furnace cooling. The cooling rate for an argon cool is 100°C per min for the first minute, with an average cooling rate of 35°C per min over a 15-min period. A furnace cool requires about 10 hr, with an average cooling rate of 3.6oC per min during the first hour, and an average cooling rate of 1.2°C per min over the 10-hr period. Microimpact Test: The specimen adopted was based on the cylindrical Izod Type Y specimen (ASTM, E23-41T). All dimensions were reduced to half scale, including the notch radius. Specifications are shown in Fig. 1. The specimen is held vertically in an adapter and broken as a cantilever beam. Impact tests were run on a constant-velocity (11.34 ft per sec) Tinius Olsen impact testing machine with a total available energy of 100 in.-lb. Tests were made to determine the correlation between this microimpact and the standard V-notch Charpy impact test. Curves showing impact energy as a function of temperature for both impact tests are plotted in Fig. 2. Transition temperatures, when they occur, are about the same for both impact tests. All three titanium-rich materials have the same conversion factor, 10. Tensile Testing: Tensile tests were conducted on Baldwin-Southwark testing machines using the 600, 2400, or 3000 1b range. Specifications for the test specimen were taken from the 1948 edition of the ASM Metals Handbook, and are shown in Fig. 1. Strain measurements were made using an SR-4 resistance gage (Type A-7) cemented to the reduced section in conjunction with a lever-type extenso-meter. Readings on the SR-4 strain indicator were

Jan 1, 1954

By Joseph L. Gillson

In this volume, which is divided into chapters on a commodity basis, many subjects inevitably have a common interest with others, or are interrelated in one way or another. No group of commodities is so outstanding in this regard as is that group of which the basic raw material is a carbonate rock. There are chapters in this book on dimension stone, on crushed stone, on chalk and whiting, on cement, on lime, on magnesite, on sodium carbonate; all of which are entirely, or partly, products of the "carbonate rocks." Rhodochrosite, witherite and strontianite are other carbonates described respectively in chapters on manganese minerals, barium minerals and strontium minerals; and travertine and cave onyx are carbonate materials used as ornamental stones. High grade marble is used for monuments and statues and as a decorative building stone. The aggregate tonnage of the basic raw material, the carbonate rock, is so enormous that it exceeds the tonnage of any metallic ore mined and quarried in the United States, and in turn is exceeded only by coal, sand and gravel, and by water in the gross tonnage produced by the mineral industry in this country. The specialties, including magnesite, sodium carbonate, and rhodochrosite, witherite and strontianite are produced from uncommon rock types of limited occurrence. They are described adequately in the chapters devoted to these subjects. This chapter is concerned with the limestones and dolomites. No other solid rocks are as important as these in the industrial life of the world, nor are any so widely known. Massive units of some carbonate formations form many of the spectacular features of some of the most rugged mountains in the world, such as the Dolomite Alps, the Canadian Rockies, some rugged peaks of the eastern flanks of the American Rockies in Colorado, Wyoming and New Mexico and in many ranges in Utah, Nevada and Arizona. These continue down the Sierra Oriental in Mexico and on around the world to the flanks of the Himalayas, in the South China Highlands, and separating Asia and Europe in the flanks of the Ural Mountains where they form many reexhumed hills called Shikhany.50a Less rugged and more nearly flat-lying carbonate rocks underlie most of the area of our great Middle Western plain, the southern part of England and the northern part of Africa. These are the common rock formations known from boyhood to a large part of the world's population. The use of lime, produced by the burning of limestone, goes back to antiquity, and mortar was used widely in Biblical times. The cement made by mixing volcanic ash with lime was a basic material of construction when Pompeii was built. Our Eastern states are dotted with abandoned limestone quarries, and the ruins of old lime kilns which were operated in Colonial days. Many limestones and dolomites are found in flat-lying strata of a thickness convenient for quarrying as a building stone, so that houses and other structures have been built with them for centuries. In order to replace the muddy and rutted dirt roads with surfaces that could stand the hard usage of his wagons and carts, man learned to crush stone and put a layer of it on his old trails. Such roads built by the Romans have been in use over much of

Jan 1, 1960

By M. D. Cooper

UNDERGRADUATES in mining engineering may be prepared for work by giving them sound instruction in the courses generally considered essential to the profession. The industry is not deeply concerned about the details of those courses. The average man in the coal industry does not wish to insist upon a rigid program. Therefore, he differs little from those in the teaching profession who evidently are not unanimous in their opinions, or all college catalogs would be alike. For the good of the profession, it is just as well that there should be differences in regard to details. It appears that students graduating in mining engineering from the accredited institutions receive similar instruction. It is taken for granted that the graduate will have a good understanding of English, mathematics, mechanics, electricity, chemistry, physics, geology, and surveying, in addition to his major courses in mining. Somewhat belatedly, industry hopes he will have had at least an introduction to the subject of labor relations, the importance of which is only too clear at present. The coal industry expects, of course, that students in mining engineering will be taught the strictly mining subjects by men who have had practical experience in the mines and who keep themselves well informed in regard to current methods. While the undergraduate is subject to the control of members of the teaching profession, industry expects him to be trained in certain ways that are not a part of his textbooks, but can be made an inseparable part of his development by the skillful supervision of his teachers. Of the desired characteristics, dependability is of the utmost importance. Probably most employers would overlook certain short-comings if the young graduate demonstrated that he was thoroughly dependable. If he always appeared at the right place at the right time with the proper equipment, he would soon be well established as a welcome member of his organization. The graduate who gets a reputation for being undependable will have little opportunity for advancement. Closely allied to dependability is loyalty. Athletic teams and social groups in college tend to develop loyalty which may well be carried over into industry. This does not mean that the graduate has to be satisfied with customary practices. The average manager is glad to see the graduate make constructive criticism as long as he demonstrates his loyalty at the same time. It is important that his loyalty keeps him alert arid ready to take helpful action for the benefit of his organization, and especially to stand with it during times of stress. With or without an introduction to labor relations in college, the graduate is expected to develop ability in this most important field. Beginning with himself, he will find it essential to deal agreeably with his immediate associates. Getting along in friendly fashion with his own small group will be a great help as his responsibilities increase and he is required to deal with larger numbers of persons. On a higher scale, his interest in his community may grow at the same time by voluntary work in any one of a great number of useful activities. Industry expects the graduate engineer to be a mature man at the time he gets his first job. Supposing that he has better than average intelligence, industry expects him to continue to grow intellectually and to fit himself for responsible jobs when they are offered to him. For this reason, employers are apt to look over his college record to see what he did that would indicate his fitness for leadership. There is interest in knowing what he did beyond the requirements. As evidence of his mental growth, it is expected that the graduate will do independent thinking; that he will not take too much for granted. When he reads a report, he should develop the ability to see whether the subject is new or whether it is just a description of an old method that has been superseded by something better. For the same reason, the graduate should be able to accept conditions that have been arrived at by sound experience rather than cling to something else that seems better in theory. In this connection, it may be remarked that the ability to operate successfully a personal budget will be noteworthy, as it may be assumed that a man who knows how to conduct his own affairs will be prepared to assume the larger responsibilities of industry. Membership in AIME will indicate to the employer that the graduate is interested in the mining industry as a whole. Therefore, it is good evidence of something more than a local outlook. Quite apart from college training and mental ability, the newly employed graduate will be expected to be willing to do hard manual labor for a time. This will give him an understanding of the actual conditions of work done by those he supervises later. He will gain their confidence and be able to see that the work is carried on in a safe and efficient manner. Part of this experience may be acquired in his summer vacations during his undergraduate career. Such work would make a favorable impression on a prospective employer, especially if the graduate showed a willingness to continue it until he was prepared for something better. To summarize, the man in authority in the coal industry will not quarrel with the professor of mining engineering over details of curriculum. He will be pleased if the school sends him graduates who possess a good foundation in the courses studied, and who may be depended upon to do their work faithfully and intelligently. Such men will be ready when the time comes to assume their places as leaders of an essential industry.

Jan 1, 1951

By M. D. Cooper

UNDERGRADUATES in mining engineering may be prepared for work by giving them sound instruction in the courses generally considered essential to the profession. The industry is not deeply concerned about the details of those courses. The average man in the coal industry does not wish to insist upon a rigid program. Therefore, he differs little from those in the teaching profession who evidently are not unanimous in their opinions, or all college catalogs would be alike. For the good of the profession, it is just as well that there should be differences in regard to details. It appears that students graduating in mining engineering from the accredited institutions receive similar instruction. It is taken for granted that the graduate will have a good understanding of English, mathematics, mechanics, electricity, chemistry, physics, geology, and surveying, in addition to his major courses in mining. Somewhat belatedly, industry hopes he will have had at least an introduction to the subject of labor relations, the importance of which is only too clear at present. The coal industry expects, of course, that students in mining engineering will be taught the strictly mining subjects by men who have had practical experience in the mines and who keep themselves well informed in regard to current methods. While the undergraduate is subject to the control of members of the teaching profession, industry expects him to be trained in certain ways that are not a part of his textbooks, but can be made an inseparable part of his development by the skillful supervision of his teachers. Of the desired characteristics, dependability is of the utmost importance. Probably most employers would overlook certain short-comings if the young graduate demonstrated that he was thoroughly dependable. If he always appeared at the right place at the right time with the proper equipment, he would soon be well established as a welcome member of his organization. The graduate who gets a reputation for being undependable will have little opportunity for advancement. Closely allied to dependability is loyalty. Athletic teams and social groups in college tend to develop loyalty which may well be carried over into industry. This does not mean that the graduate has to be satisfied with customary practices. The average manager is glad to see the graduate make constructive criticism as long as he demonstrates his loyalty at the same time. It is important that his loyalty keeps him alert arid ready to take helpful action for the benefit of his organization, and especially to stand with it during times of stress. With or without an introduction to labor relations in college, the graduate is expected to develop ability in this most important field. Beginning with himself, he will find it essential to deal agreeably with his immediate associates. Getting along in friendly fashion with his own small group will be a great help as his responsibilities increase and he is required to deal with larger numbers of persons. On a higher scale, his interest in his community may grow at the same time by voluntary work in any one of a great number of useful activities. Industry expects the graduate engineer to be a mature man at the time he gets his first job. Supposing that he has better than average intelligence, industry expects him to continue to grow intellectually and to fit himself for responsible jobs when they are offered to him. For this reason, employers are apt to look over his college record to see what he did that would indicate his fitness for leadership. There is interest in knowing what he did beyond the requirements. As evidence of his mental growth, it is expected that the graduate will do independent thinking; that he will not take too much for granted. When he reads a report, he should develop the ability to see whether the subject is new or whether it is just a description of an old method that has been superseded by something better. For the same reason, the graduate should be able to accept conditions that have been arrived at by sound experience rather than cling to something else that seems better in theory. In this connection, it may be remarked that the ability to operate successfully a personal budget will be noteworthy, as it may be assumed that a man who knows how to conduct his own affairs will be prepared to assume the larger responsibilities of industry. Membership in AIME will indicate to the employer that the graduate is interested in the mining industry as a whole. Therefore, it is good evidence of something more than a local outlook. Quite apart from college training and mental ability, the newly employed graduate will be expected to be willing to do hard manual labor for a time. This will give him an understanding of the actual conditions of work done by those he supervises later. He will gain their confidence and be able to see that the work is carried on in a safe and efficient manner. Part of this experience may be acquired in his summer vacations during his undergraduate career. Such work would make a favorable impression on a prospective employer, especially if the graduate showed a willingness to continue it until he was prepared for something better. To summarize, the man in authority in the coal industry will not quarrel with the professor of mining engineering over details of curriculum. He will be pleased if the school sends him graduates who possess a good foundation in the courses studied, and who may be depended upon to do their work faithfully and intelligently. Such men will be ready when the time comes to assume their places as leaders of an essential industry.

Jan 1, 1951

By D. C. Hilty

It has long been known that in melting high-chromium steels, some of the carbon might be oxidized out of the melt without excessive simultaneous oxidation of chromium, and that higher temperatures favor retention of chromium. The advent of oxygen injection as a tool for rapid decarburization of a steel bath permits significantly higher bath temperatures, and it was quickly recognized that the use of oxygen injection facilitated the oxidation of carbon to low levels in the presence of relatively high residual chromium contents. Up to the present time, however, specific data pertaining to the chro-mium-carbon-temperature relations in chromium steel refining have not been available. Individual steelmakers have evolved practices more or less empirically, but there has been very little real basis for predicting how effective any given practice can be in permitting maximum oxidation of carbon with minimum loss of chromium. The current investigation, therefore, was undertaken in an effort to establish the fundamental carbon-chromium relationship in molten iron under oxidizing conditions. As reported below, the equilibrium constant and the influence of temperature on that constant have been derived for the iron-chromium-carbon-oxygen reaction in the range of chromium steel compositions with what appears to be a fair degree of precision. The practical application of the result will be obvious. Experimental Procedure The laboratory investigation was carried out on chromium steel heats melted in a magnesia crucible in a 100-lb capacity induction furnace at the Union Carbide and Carbon Re- search Laboratories. The charges for the heats consisted of Armco iron, low-carbon chromium metal, and high-carbon chromium metal, the relative proportions of which were calculated so that the various heats would contain from approximately 0.06 pct carbon and 8 pct chromium to 0.40 pct carbon and 30 pct chromium at melt-down. When the charges were melted, the bath temperatures were raised to the desired level, and the heats were then decarburized by successive injections of oxygen at the slag-metal interface through a ½-in. diam silica tube at a pressure of 30 psi. The duration of the oxygen injections was from 30 sec to 2 min. at intervals of approximately 5 to 30 min. It did not appear that length or frequency of the injection periods had any significant effect on the results; cansequently, no effort was made to hold them constant and they were controlled only as was expedient to the general working of the heats. Between successive injections, the heats were sampled by means of a copper suction-tube sampler that yields a sound, rapidly-solidified sample representative of the composition of the molten metal at the temperature of sampling. This sampling device is a modification of the one described by Taylor and Chipman.1 An attempt was made to vary bath temperatures between samples, but it quickly became evident that, unless the variations were small or unless the new temperature was maintained for a minimum of 15 min. during which an injection of oxygen was made in order to accelerate the reactions, a very wide departure from equilibrium resulted. For most of the runs, therefore, temperature was maintained relatively constant at approximately 1750 or 1820°C. A few reliable observations at other temperatures, however, were obtained. Temperature Measurement The high temperatures involved in this investigation were measured by the radiation method, utilizing a Ray-O-Tube focused on the closed end of a refractory tube immersed in the metal bath. The immersion tubes employed were high-purity alumina tubes specially prepared by the Tona-wanda Laboratory of The Linde Air Products Co. These tubes were quite sturdy under reasonable mechanical stress at high temperature. They were unusually resistant to thermal shock, and chemical attack on them by the melts was slow. With care, it was found possible to keep these tubes continuously immersed in a heat for as long as 5 hr at temperatures up to 1850°C, before failure by fluxing occurred. The Ray-O-Tube—alumina tube assemblage was similar to those supplied commercially for lower temperature applications. In operation, the alumina tube was slowly immersed in the molten metal to a depth of approximately 5 in., and the device was then clamped solidly to a supporting jig where it remained for the duration of the run. A photograph of the equipment, in operation with Ray-O-Tube in place and oxygen injection in progress, is shown in Fig 1. When in position in a heat, the instrument was calibrated by means of an immersion thermocouple and an optical pyrometer. For calibration through the range of temperatures from 1500 to 1650°C, a platinum -platinum + 10 pct rhodium thermocouple in a silica tube was immersed alongside the alumina tube. Output of the Ray-O-Tube in millivolts and the

Jan 1, 1950

By Charles L. Mantell, George Ferment

This investigation determined the jeasibility of current-potential curves as an analytical tool for monitoving manganese electroicinning solutions for metallic impurities. Nine metallic impurities were studied: trickel, cobalt, silver, copper, zinc, molybdenum, catlmiun magnesium, and sodium. The individual effect of each metallic irzplrl,iLy on the standard point-izatiorr was determined for a range of 'concentration. Not all of Ihe impurities affected the polarization curve; correlation between transition current and cotrcetztration of impurity. Molybdenum had an uncorrelated effect on the polarization curie. The effects of binary mixtures of impurities were studied to see if the transition current of the mixture could be predicted. Interactions occurred between impurities which prei'etzted the pvecliction the transition current of tile rrlisLitr.e by additive laws. The polarization curve was shown to be sensitive to metallic impurities which affect the current efficiency of a manganese cell. THE application of current-potential curves in manganese electrowinning is not new but played an important part in the conversion of an inoperative large-scale pilot plant into a continuous comnercial operation.1"" The effects of trace metallic impurities in manganese electrowinning solutions on cathode current-potential curves,5-19 specifically nine impurities (nickel, cobalt, silver, copper, zinc, molybdenum, cadmium, magnesium, and sodium) were studied for a range of concentrations. Cadmium, magnesium, and sodium had no effect. Nickel, cobalt, silver, copper, and zinc yielded a correlation between transition current and concentration of impurity: molybdenum had an uncorrelated effect. Binary mixtures of impurities were studied to de-termine if the transition current could be predicted from the values for the individual constituents. Interactions between impurities prevented the prediction of the transition current of the mixtures by an additive law. The polarization curve was shown to be sensitive to metallic impurities which affected the rurrent efficiency of a manganese cell. The electrowinning of manganese relies upon maintaining a high-purity electrolyte. Purification schemes were developed to remove all of the heavy metals, the magnesium and calcium. Concentration limits have been established for single impurities. With high-purity electrolytes, favorable cell efficiencies are consistently obtained; but with impure solutions there is a reduction in efficiency. With the rapid rise to commercial prominence of electrochemical processes, there has been a lag in developing rapid methods for determining whether an electrolyte is sufficiently pure for efficient electrolysis. The most sensitive test for impurities is the behavior of the electrochemical cell. Impure electrolytes will cause a serious reduction in cell efficiency, and in many cases metal deposition will be prevented or reversed. The effects of impurities are not usually evident until after long periods of electrolysis, resulting in a disruption of the process. A method which could predict the behavior of an electrolyte in terms of cell efficiency prior to reaching the commercial cell is needed. Van Arsdale and Maier," Allmand and Campbell,a1'22 and others attempted to develop an electrochemical process for manganese,"3"29 but all met with some degree of difficulty. Jacobs et a studied the effects of metallic impurities in manganese electrowinning. Plant methods were based on a 2-hr plating in a small electrolytic cell. The final method was based on a 24-hr plating run, performed in a relatively large test cell. Here, current efficiencies were calculated with and without impurities in the electrolyte, and the critical concentration was established as the maximum amount of impurity that the electrolyte could tolerate before there was a serious loss in efficiency. The amount of impurity that could be tolerated decreased as the plating time increased, indicating deposition of the impurity. This test duplicated the conditions of commercial practire, and was sensitive to impurities. Small amounts of impurities significantly affect the hydrogen overvoltage at a metal surface, increasing or decreasing it depending upon the nature of the impurity. ockris established the limit for the onset of poisoning at 10-l moles per liter. The effects of impurities were studied by potential-time curves at a constant current density. The technique did not lend itself to rapid analysis in that several hours were needed to establish stable potential-time curves. Serfass and coworkers3' developed calorimetric and spectrophotometric techniques for analyzing for trace impurities. Our method4'" is based on current-cathode potential curves produced by continuously changing the cell voltage at a programmed rate. Starting with a clean stainless-steel cathode, the cell voltage was initially adjusted to yield zero current: the program was then started and the current was plotted as a function of the cathode potential. Fig. 1 is a typical current-potential curve. In initial region A the hydrogen-evolution reaction predominates and the polarization curve

Jan 1, 1967

By S. J. Sindeband

Prior to discussing the metallurgy of sintered chromium borides, it is pertinent to outline some of the reasoning behind this investigation and the purposes underlying the work. This study was initiated as an aproach to the ubiquitous problem of a material for service at high temperatures under oxidizing atmospheres, and it was undertaken with a view to raising the 1500°F (816°C) ceiling to 2000°F (1093°C) or better. For the reason that no small, but rather a major, lifting of the high temperature working limit was being attempted, it was felt appropriate that a completely new approach be taken to this problem. A summary of the thinking behind this approach was published recently by Schwarzkopf.' In briefest terms, it was postulated that the following requirements could be set up for a material which would have high strength at high temperatures. 1. The individual crystals of the material must exhibit high strength interatomic bonds. This automatically leads to consideration of highly refractory materials, since their high energy requirements for melting are related to the strength of their atom-to-atom bonds. 2. On the polycrystalline basis, high boundary strength, superimposed on the above consideration, would also be a necessity. Since this implies control of boundary conditions, the powder metallurgy approach would hold considerable promise. Such materials actually had been fabricated for a number of years, and the cemented carbide is the best example of these. Here a highly refractory crystal was carefully bonded and resulted in a material of extremely high strength. That this strength was maintained at high temperature is exhibited by the ability of the cemented carbide tool to hold an edge for extended periods of heavy service. Nowick and Machlin2,3 have analytically approached the problem of creep and stress-rupture properties at high temperature and developed procedures whereby these properties can be approximately predicted from the room temperature physical constants of a material. The most important single constant in the provision of high temperature strength and creep resistance is shown to be the Modulus of Rigidity. On this basis, they proposed that a fertile field for investigation would be that of materials similar to cemented carbides, which have Moduli of Rigidity that are among the highest recorded. The cemented carbide, however, does not have good corrosion resistance in oxidizing atmospheres and without protection could not be used in gas turbines and similar pieces of equipment. It would be necessary then to attempt the fabrication of an allied material based upon a hard crystal which had good corrosion resistance as well. It was upon these premises that the subject study was undertaken and at an early stage it was sponsored by the U.S. Navy, Office of Naval Research. Since then, it has been carried on under contract with this agency. Chromium boride provided a logical starting point for such research, since it was relatively hard, exhibited good corrosion resistance, and, in addition, was commercially available, since it had found application in hard-surfacing alloys with iron and nickel. That chromium boride did not provide a material that met the ultimate aim of the study results from factors which are subsequently discussed. This, however, does not detract from the basis on which the study was conceived, nor from the value of reporting the results which follow. Chromium Boride While work on chromium boride proper dates back to Moissan,4 there has been a dearth of literature on borides since 1906. Subsequent to Moissan, principal investigators of chromium boride were Tucker and Moody,5 Wede-kind and Fetzer,6 du Jassoneix,7,8,9 and Andrieux." These investigators were generally limited to studies of methods of producing chromium boride and detennining its properties. Some study, however, was devoted to the chromium-boron system by du Jassoneix,7 who did this chemically and metal-lographically. This system is not amenable to normal methods of analysis by virtue of the refractory nature of the alloys involved, and the difficulties of measurement and control of temperature conditions in their range. Dilatometric apparatus is nonexistent for operation at these temperatures. Du Jassoneix made use of apparent chemical differences between two phases observed under the microscope and reported the existence of two definite compounds, namely: Cr3B2 and CrB. These two compounds, he reported, had quite similar chemical characteristics, but were sufficiently different to enable him to separate them. The easiest method for producing chromium boride is apparently the thermite process, first applied by Wede-

Jan 1, 1950

By W. A. Hockings, M. E. Volin, A. L. Mular

Batch grinding tests at short times were made in a laboratory rod mill with 10 x 14 mesh quartzite in corn syrup-water mixtures of varying viscosity. The weight fraction broken and size modulus were found to be independent of viscosity up to 20 cp, but at higher viscosities breakage of the feed decreased while the size modulus increased. The mechanisms of the effect of viscosity are discussed on the basis of particle dynamics. The effect of viscosity of the suspending medium on grinding has not been accorded much attention, although both wet and dry grinding are used in bene-ficiation processes and the viscosities of water and air differ greatly. schweyerl investigated the rates of grinding of quartzite in a pebble mill with air, water and glycerol as the suspending media and found the rate in terms of new surface developed per thousand revolutions of the mill to be constant and dependent on viscosity up to 20,000 rev, but to decrease and become independent of viscosity for longer grinding times. Based on considerations of the drag forces exerted on suspended particles and on the grinding media under turbulent and laminar flow conditions, viscosity can be expected to inhibit the rate of grinding in a given system as it changes the flow conditions from turbulent to laminar. The objective of this study was to determine the effect of viscosity of the suspending fluid in the batch grinding of a homogeneous feed of uniform size for short residence times, a region of more practical interest than the long times investigated by Schweyer. In actual grinding operations, the viscosity of the fluid does not change, but the consistency (apparent viscosity) of the pulp increases with additions of fines. In this study the consistency of the pulp was not measured because it was thought that the grinding times were too short to alter the apparent viscosity appreciably.* *Minus 200 mesh quartz at 65% solids has an apparent viscosity of about 25 cp as measured by a simple consistometer. Personal communication from D. F. Kelsall to A. L. Mular. For the purposes of this study, the effect of viscosity on grinding can be shown sufficiently well on the basis of the cumulative weight fraction finer than the feed size at any time, and the size modulus, ku, of the finer than feed sizes. The parameter, ku, is obtained from a form of the Gaudin-Schuhmann equation. METHOD AND MATERIAL The grinding tests were accomplished in a laboratory rod mill 10-1/2 in. long and 8 in. bore with a 20-lb rod charge consisting of two 1-in., eight 3/4-in., eight 1/2-in., and eight 1/4-in. diam rods. The mill speed was 43.2 rpm. Charges of 400 g 10 x 14 mesh Wisconsin quartzite** were ground for times of 60, **Courtesy of Minnesota Mining and Mfg. Co. 120, 180 and 240 sec in 500 cc of fluid made up of corn syrup and water in measured proportions. The temperature of the pulp was measured immediately upon completion of grinding, and the size distribution of the quartzite was determined by a wash-wet-dry screening technique. Corn syrup was chosen because of its high viscosity and ideal viscous behavior, and also because its density is not greatly different from that of water. The viscosity of each fluid mixture was measured with an Ostwald viscometer at 25 C. and a correction was made for the temperature of the pulp according to tabulated values for sucrose solutions. The density differences were so small that they probably did not significantly affect breakage. RESULTS AND DISCUSSION The conditions of the grinding tests, computed viscosities, weight fractions finer than the feed size and values of a and ku are listed in Table I. The values of a and ku were taken from the linear portions of the log-log plots of cumulative weight fraction finer vs size for the various grinding times. The plots for tests in fluids of average viscosity 8.6 cp are shown in Fig. 1. In general all of the sets of curves were linear throughout most of the size

Jan 1, 1965

By R. Bainbridge

THE Consolidated Mining and Smelting CO. of Canada Ltd. has operated a lead smelter at Trail, B. C., for many years. In order to take advantage of metallurgical advances, as well as to improve materials handling methods, this company, commonly known as "Cominco," commenced planning a program of smelter revision and modernization some years ago. The first stage of this program involved the design and construction of a new blast furnace gas cleaning system. The selection of equipment, the design of facilities, and preliminary operating details of this system will be dealt with in this paper. The essential problem was to clean and collect 100 tons of dust daily from 153,000 cfm* (12,225 lb per min) of lead blast furnace gas which varied in temperature from 350º to 1100°F. Because it was desired to collect the dust dry, either a Cottrell or a baghouse cleaning plant was to be selected. Comin-co's many years of experience with both systems provided a background for choosing the most satisfactory installation. All information pertinent to the two methods of dust recovery was carefully investigated, and it was decided to replace the existing equipment with a baghouse. Very briefly, the reasons for this decision were as follows: 1—A baghouse installation would be practical because the SO2 content of the gas was low and corrosion would not be a problem if the baghouse operating temperatures were held sufficiently above the dew point. 2—Variations in the physical characteristics of fume and dust, which are inherent in this blast furnace operation, should not substantially affect the operating efficiency of a baghouse. 3—For the same capital cost, metal losses (stack and water losses) would be appreciably less in a baghouse. 4—A baghouse would be easier to operate, and would not require the use of highly skilled labor. 5—Operating and maintenance costs of a bag-house would be lower. 6—The only available space for reconstruction was relatively small, and not suited to a Cottrell installation. Once the baghouse system was decided upon, detailed design of the installation was begun. Baghouse Design Gas Cooling: Before the required capacity of the baghouse could be determined, the method of cooling the gas to the temperature necessary for bag-house operation had to be chosen. The problem confronting the design engineers was how best to cool 153,000 cfm of gas from a temperature ranging from 350°F to brief peaks of 1100°F, down to 210°F, the maximum safe baghouse inlet temperature. A survey of existing blast furnace gas temperatures in the outlet flue showed that the normal range was as given in Table I. The obvious choices of cooling method were: 1— cool completely by the addition of tempering air; 2—utilize a heat exchanger; 3—cool by radiation; and 4—cool with water spray in conjunction with the admission of tempering air. The advantages and disadvantages of the various cooling methods were: Air Addition: To cool completely by the admission of tempering air involved tremendous volumes, Fig. 1. For example, to cool 1 lb of blast furnace gas at 450°F requires 1.84 lb of air at 80°F or 1.60 lb at 60°F. As it is necessary to design for peak conditions, it can readily be seen that volumes of tempering air in the order of 1,500,000 cfm would have to be handled. Using the normal design figure of 2.5 cu ft per sq ft of bag area, a baghouse installation comprising some 600,000 sq ft of filter cloth would be necessary. Such design requirements would be prohibitive, not only from a standpoint of capital expenditure, but also because of space limitations. Heat Exchanger: The utilization of a heat exchanger was given serious consideration. A horizontal tube unit using air as the medium to cool the required volume of blast furnace gas from 400" to 250°F was investigated. Cooling above 400°F would be done by water spray, and below 250°F by admission of tempering air. The estimated capital cost of such a unit was found to be prohibitive. From an operating standpoint, there was considerable doubt as to whether the soot blowing equipment provided would effectively keep the dust from building up on the tube surface. The performance of heat exchangers operating on dusty gas in other company operations had not been too favorable. Radiation Cooling: Although somewhat cumbersome, gas cooling by radiation from 'trombone' tubes or other similar equipment (cyclones) is employed in many metallurgical operations. Such an installation was also considered. However, calculations showed that an installation much larger than the space available would be required to handle the gas volume involved. For example, to cool 153,000 cfm of blast furnace gas from, say, 600' to 250°F (i.e., remove in the order of 58,500,000 Btu per hr with heat transfer rates varying from 1.1 Btu per sq ft per hr per OF for the higher temperature ranges to 0.88 Btu per sq ft per hr per OF for the lower ranges) would need a cooling area of some 175,000 sq ft.

Jan 1, 1953

By Eugene F. Broome, Hugh A. Walls

A diffusion measurement technique based on a shear cell comprised of only two segments is described. The diffusion boundary value problem for the finite capillary geometry is solved in general for any arbitrary initial concentration profile and is subsequently specialized for the modified shear cell problem. Effects of convection and mixing at the shear interface were found to be negligible. Mercury self-diffusion coefficients were determined from -25° to 252°C. These data are in good agreement with those found by Meyer. ALTHOUGH diffusion in liquid metals has been of interest for over two centuries, the need for measurement techniques of improved accuracy and precision has become increasingly apparent as additional data have been obtained and theory has become more refined. These conditions reflect the experimental difficulties inherent in liquid diffusion measurements, in which transport by other processes, such as convection, tends to mask the diffusive transport. Frequently the disagreement between several theoretical predictions is less than that found between different sets of data obtained for a system. Moreover, as has been shown by Nachtrieb,1 diffusion data are needed over much larger temperature ranges if the functional dependence on temperature is to be known. Thus, improved techniques must be devised if experimental data are to augment fundamental understanding of the liquid state and to meet technological needs. The available techniques have been discussed elsewhere.' Of these, only the capillary-reservoir, long capillary, and shear cell techniques will be discussed briefly in terms of experimental advantages and disadvantages. These methods served to establish design criteria for the modified shear cell described here. The capillary-reservoir technique of Anderson and saddington3 has been the most widely used method in recent years. The method offers experimental simplicity relative to other methods and has been employed for high-temperature measurements. Moreover, the mathematical relationship between the measured concentration ratio and the diffusion coefficient is such that smaller values of the ratio are achieved for a specified diffusion time relative to other methods. The amplified errors between the concentration ratio and the calculated diffusion coefficient are diminished at lower values of the ratio.' The method also permits multiple determination by the simultaneous use of several capillaries. Disadvantages of the capillary-reservoir method are primarily associated with the hydrodynamic ef- fects of convection and of placing the capillary in the reservoir. These effects are most pronounced in the region near the open end of the capillary and produce an ill-defined boundary condition between the capillary and the reservoir. Such effects are not amenable to experimental or mathematical correction2 (although this has been suggested4). The long-capillary method of Careri, Paoletti, et al.5-10 involves filling one half of a small capillary tube of 150 to 200 mm total length with material of one composition or radioactivity and the other half with the second part of the diffusion couple. This arrangement eliminates the adverse hydrodynamic effects associated with the capillary-reservoir technique; however, certain other experimental difficulties are encountered in this method. The more significant of these difficulties involve the melting, expansion, contraction, and solidification of the diffusion system. The dependence in some cases of the diffusion coefficient on the capillary diameter noted by Careri et a1.7 (termed the "wall effect") has been alternatively explained by Nachtriebl as a convection effect during solidification. In mutual diffusion measurements, the convection problems associated with melting and solidification are increased because of the differences in melting points and in expansion coefficients between the halves of the diffusion couple. However, the errors caused by convection effects within this method are usually less than those in the capillary-reservoir method. Furthermore, the concentration profile needed to determine concentration-dependent diffusion coefficients by the Boltzmann-Matano analysis can be obtained from this method. Of the previous attempts to use shear cells, only the cell used by Nachtrieb and Petit11,12 appears to have yielded good data. They reduced the mechanical complexity of the conventional shear cell by using a cell comprised of only four segments. Three of these segments were filled with ordinary mercury and the fourth with radioisotopic mercury in their determination of mercury self-diffusion coefficients. The average concentration (radioactivity) was determined in each segment following a period of isothermal diffusion. These concentration values were fitted to concentration profiles obtained from the Stefan-Kawalki tables, and the diffusion coefficients were evaluated. Thus, although the number of cell segments is reduced in their method, some information about the concentration profile can be obtained in terms of the Stefan-Kawalki analysis. Moreover, their cell is suitable for measurement of diffusion coefficients at elevated pressure, as they successfully demonstrated with mercury. Consideration of the design and experimental features of the methods discussed above suggested several criteria for the new cell: 1) a ''total" capillary system, as opposed to a capillary-reservoir system, should reduce adverse convection effects; 2) such a capillary system should avoid the problems en-

Jan 1, 1969

By H. C. Theuerer

GeC14 may be purified by extraction with HCI and chlorine. The process is as effective for the removal of AsCI:, as the more cumbersome distillation methods usually used for this purpose. GERMANIUM for semiconductor use contains impurities at levels no higher than a few parts in ten million. Material of this quality is obtained from highly purified GeC1, by hydrolysis to the oxide and reduction of the oxide in hydrogen. When purifying GeCl,, AsC1, is the most difficult impurity to remove. This is usually accomplished by multiple distillation procedures.1-3 AsC1, may also be removed from GeC1, by extraction with HC1.1-4 Reducing the arsenic to low concentrations is not practicable, however, because of the large number of extractions needed. This paper discusses a new method for the removal of arsenic from GeC1, by extraction with HC1 and chlorine. The method is rapid, leads to little loss of germanium and is at least as efficient as the distillation procedures currently being used. Theory of Extraction Procedures In the simple extraction of GeC1, with HC1, the following reaction occurs ASCl8G8C1 D AsCl3rc1 at equilibrium CA/Cn= K, where K is the distribution coefficient, and C, and C,, are the molar concentrations of AsC13 in HC1 and GeCl,, respectively. The materials balance equation for this reaction is VACA + vncn = VnC,, where V, and Vn are the volumes of HC1 and GeCl4, respectively, and C, is the initial concentration of AsC13 in GeC1,. From this it can be shown that for multiple extractions where C,, is the concentration of AsC13 in GeC14 after n extractions, and r is the ratio of V, to V,,. It is assumed that r is maintained constant, that equilibrium is established during each extraction, and that K is independent of the AsCl3 concentration. By saturating the system with chlorine, the following reaction occurs in the aqueous phase AsCl3 + 4H2O + Cl2 D H5AsO4 + 5HC1 at equilibrium K' = ------------ ai - a4 h2u - aet2 where a is the activity of the various components. The effect of this reaction is to reduce the concentration of the AsC1, in the aqueous layer and, therefore, to promote further extraction of this component from the GeC1, layer. If the arsenic acid remains entirely in the aqueous phase, the net effect of this reaction is to promote the removal of arsenic from the GeC11. The materials balance equation for extraction with HC1 and chlorine with the foregoing reaction is, then, VaCC + VACA + VACn = VnCo where C,. is the molar concentration of H3AsO, in the HC1. With the added assumptions that the activities of AsC13 and H8ASO4 in the aqueous phase are equal to their molar concentrations, it can be shown that for n extractions Cn/Cu = (1/rkK + rK + 1) n where k - K1 a4h2o - acl2/aoncl. It can be seen by comparing Eqs. 1 and 2 that if k is large, the removal of AsC1, by HC1 extraction will be greatly improved by the addition of chlorine. Dilution of the HCI used in the extraction with chlorine would also favor the separation. This, however, would increase the loss of GeCl,, which is undesirable. Experimental Procedure Germanium prepared from oxide of semiconductor purity is n-type with resistivities greater than one ohm-cm. The resistivity is controlled by the donor concentration, which is —lo-: mol pct. Germanium prepared from material with added arsenic will have lower resistivity commensurate with the arsenic concentration. With such material, at arsenic concentrations above 10-1 mol pct the resistivity is controlled by the added arsenic, and effects due to other impurities initially in the oxide are negligible. In this investigation GeO, of semiconductor purity was converted to GeCl,, and -0.01 mol pct As was added. This material was used for the extraction experiments and the purification attained determined by a comparison of the resistivity data for samples of germanium prepared from the initial and purified GeC1,. A method for calculating the arsenic concentration from the resistivity data is discussed later. The details of the experimental procedures used are as follows: Two hundred and thirty cu cm GeC1, were prepared by the solution of GeO, in HC1, followed by

Jan 1, 1957

By R. S. Busk

Niels Engel (University of Alabama, University, Ala.)— In this paper it was pointed out that the electron-gas and energy-band theory accounts for the fact that the lattice parameters exhibit a sudden change when the electron concentration (number of bonding electrons per atom) exceeds a certain number around two. This statement is said to support and prove the electron-gas theory. But this theory is not able to account for a series of experimental data. Also several expectations, deduced from this theory, are not found to exist. In Figs. 6 and 7 the energy bands of the second and third periods are given as they must be assumed in order to account for the electrical properties of the elements in these periods. In Figs. 6 and 7 the electron-gas and energy-band theory is compared with the electron-oscillator hypothesis in accounting for the properties of the elements in the second and third periods. Fig. 6 shows the second period, The energy-bands are overlapping and separated to be in agreement with the electrical conductivity of the elements. The oscillator hypothesis explains conductivity due to electron vacancies. In graphite there is a closed s-shell in every other atom and two vacancies in the others. Conductivity is therefore only maintained by migration of s-electrons in graphite. In boron there are no s-electrons. The diatomic molecules of nitrogen and oxygen and the paramagnetism of oxygen can be accounted for by a similar behavior as the s-electrons of the bonding electrons. But this explanation will deviate too much for the purpose of this discussion. Fig. 7 shows the third period. In the energy-band picture about two s-electrons are assumed in magnesium and aluminum, but only one s-electron is assumed in silicon. The diamond lattice is assumed to be controlled by a sp3 hybrid. However the electron distribution develops ideally according to the oscillator hypothesis. Only sodium, magnesium, and aluminum exhibit electron vacancies and conductivity. To account for the insulator properties in Si, P, and S in the third period it must be assumed that the four last added p-electrons must be taken up in bands containing only one electron per band.' (Compare the electron band picture in Hume-Rothery.' Hume-Rothery does not consider the insulator properties of the nonmetals.) In the second period already the first p-electron must have entered a single electron band. Based on the energy-band picture in Figs. 6 and 7, the following questions must be asked: 1—Is it consistent with the energy-band idea that electrons of the same kind (p-electrons) can be divided into separated bands? 2—Is it consistent with the energy band idea that single electron bands can exist? 3—Why are the first two p-electrons (in boron and diamond) separated into two single electron bands in the second period, but overlapping in the third period (aluminum)? 4—Why are s-electrons and d-electrons taken up in continuous overlapping bands, while p-electrons are divided into single electron bands? 5—Why do the peaks and valleys (y and w and further x and z) of the energy band below four electrons per atom not show up in the electrical conductivity of alloys? For example consider the Li-Mg system or the alloys between Mg and three electron metals where the mentioned discontinuity in the lattice parameter is found. 6—Why does the beginning of the p-electron band (x) not show up in the lattice constants similar to the filling up of the s-electron band (z) ? In magnesium alloys the electron-gas theory postulates the first Brillouin zone to be filled at about two electrons per atom. This is claimed to explain the sudden change in lattice spacing and c/a values of several magnesium alloys when the electron concentration exceeds a few percentage points over two electrans per atom. This was emphasized in the paper by Busk. If the electron-gas energy-band theory is correct a sudden change in electrical conductivity and possibly other properties .should be expected when the same electron-concentration or temperature is exceeded. A sudden change in lattice spacing or other properties should also be expected when the filling degree is such that p-electrons are introduced into the p-band, for example at x in Figs. 6 and 7. Such phenomena are at found by experiment. and If the number of electrons should vary with the energy level depending on the average number of bonding electrons per atom, the electrical conductivity should be expected to vary in accordance with the energy band layout (Figs. 6 and 7) caused by different numbers of conducting electrons at different filling up degrees. Nothing indicating such a behavior is observed. In addition to these discrepancies between the electron-gas and energy-band theory and measured data, the theory violates the principles developed along with the Bohr theory of atomic structure. According to these principles a filled shell is saturated and therefore unable to form bonds. Therefore two S-electrons per atom should form a closed or saturated shell, which has been pointed out as accounting for the inability of helium to form bonds. Beryllium, magnesium, or calcium atoms with two s-electrons should be expected to form inert atoms with properties almost like the helium atoms. Several other inconsistencies and disagreements with measured data of the energy-band theory can be mentioned. Some of these are discussed with reference to other papers. 8 Because the electron-gas and energy-band theory seems to fail on several points, I have developed another theory which can account for all the phenomena the electron-gas theory is able to account for. This new theory is further able to account for things which are impossible to explain by the electron-gas theory at the present state.

Jan 1, 1953

By G. H. Anderson

H. R. HANLEY*—I have been asked to discuss briefly the development of rotating cathodes for the electrolytic deposition of cadmium. The earliest recorded use of rotating cathodes was by Hoepfner at Frufurt, Germany about sixty years ago. He elec-trolized zinc chloride solution using diaphragms to separate electrodes. In the early experimental work of the Bully Hill Copper Mining and Smelting Co., Shasta County, Calif., rotating aluminum cathodes 4 ft in diam were used in the electrolysis of an acid zinc sulphate solution. Finished cathodes weighing up to 400 lb were produced. Because of mechanical difficulties, this type of cathode was abandoned for zinc, but was later used for cadmium because of the relative smoothness of deposit in comparison with stationary plates with comparable current densities. Cadmium sponge which forms on the cathode at moderate current densities (without special treatment) is entirely eliminated by a slow rotation. The rate of rotation of the cathode has an effect on the mechanical nature of the deposit. A high rate of rotation concentrates the adhering electrolyte on the shaft; a moderate rate appears to concentrate on the cathode a short distance out from the shaft tending to corrode the deposit in the form of a ring. At a very slow rotation (2 to 3 rpm) the adhering electrolyte gravitates nearly vertically, thus avoiding the cutting ring referred to above. The true explanation for the smoother deposits obtained on rotating cathodes may not be given definitely as the numerous factors involved are not thoroughly understood. Smooth deposits are obtained when the orderly growth of the metal crystals in the cathode lattice are disorganized. Thus the crystals form and grow for a very short interval when they are arrested and a new crystal forms. The continued growth of the original crystals provides large crystals and a rough deposit. Also if the acidity of the electrolyte is low, hydrogen gas bubbles adhere to the deposit. As the cathode is rotated the gas surface is brought into the atmosphere where they burst; thus the deposit is made on a surface relatively gas-free. An aluminum hub distance piece was riveted to each aluminum disk 4 ft in diam, slipped on a 4 1/2 in. steel shaft and pressed tight to prevent acid electrolyte seeping through to the shaft. The 9-cathode assembly was supported on insulated bearings. Electrical contact to the shaft was made through what was equivalent to a copper pulley. Sufficiently high conductivity brushes were placed on the face of the pulley to lead the current to the cathode bus bar. The assembly was driven by a link belt contacting a sprocket insulated from the shaft. The lead anodes were semicircular and supported on porcelain insulators placed on the bottom of the cell. Two anodes were provided for each cathode to permit an 8-in. space between them without increasing the ohmic resistance. This ample spacing permitted easy stripping of deposit with the assembly in place. Cathode cadmium was melted under 650 W cylinder oil. After casting, the primary slabs were remelted under molten caustic soda and cast into pencils 1 1/32 in. in diam. Rotating cathodes for deposition of cadmium are used at Risdon, Tasmania, and at Magdeburg, Germany. W. G. WOOLF*—This paper is very-interesting to me because in our work at the Electrolytic Zinc Plant of the Sullivan Mining Co. we had an exactly similar problem—that is, a method of producing cadmium from our purification residue, the recovery of the contained copper as a copper precipitate which could be sent to a copper smelter and the production of merchantable cadmium. It is interesting to me, not knowing of the work of the Risdon people, how closely we approximate them in their main metallurgy, diverging at several interesting steps which I would like to discuss for just a moment. For example, at Risdon they oxidize their purification residue. In our practice we take the current residue as it is produced in the purification department of the zinc plant and process it in the cadmium plant. The only oxidation that it obtains is the oxidation in the presses, the dumping of the presses and the collection and transportation of the residue to the cadmium plant. We find that the leaching of that residue does not necessarily require the oxidation step that the Risdon people evidently find necessary. The discussion of oxidation comes in again in the matter of the treatment of the precipitated cadmium sponge with zinc dust which again at Risdon is oxidized but which we do not attempt to oxidize except as it oxidizes itself in the storage. There is a partial oxidation which cannot be avoided, as Mr. David-sou pointed out, but we make no attempt to attain a complete oxidation and we dissolve the cadmium sponge in the sul-

Jan 1, 1950

By W. C. Clarke

An empirical study of the nature of the embrittle-ment which occurs in martensitic and semiaustenitic precipitation hardening stainless steels upon exposure at temperatures of from about 550" to 875°F has been undertaken. This work was aimed at determining cazusation and means of controlling this phenomenon. A commercial copper bearing precipitation hardening alloy was used as a basic material for study. The effect of heat treatment variables was studied as was the effect of variations in analysis. It is concluded from the evidence that martensitic and semiaustenitic precipitation hardening stainless steels are subject to 885°F ernbrittlement similar to that observed in the straight chromium stainless steels. A characteristic of precipitation hardening stainless steels which has limited their use in certain applications is that they embrittle when held at temperatures in the range of from about 550" to 900°F. This is true to a more or less degree in all currently available alloys, either the basically martensitic type or the semiaustenitic type. The rate of embrittlement varies markedly with exposure temperature, being low at 550" to 600°F and in-creasing as the temperature increases. In spite of this embrittlement, these alloys with their unique combination of high mechanical strength, reasonable toughness, and good corrosion resistance are used in many hundreds of applications. Nevertheless, there are potential applications where the embrittle-ment described is a limiting factor. The purpose of this investigation was to study the embrittlement of these alloys and to find a way to control or eliminate it. GUIDELINES USED IN PRESENT WORK The work reported in this paper is based on a study of 17-4 PH*, a very widely used alloy. It has been used *Trademark of Armco Steel Corp., licensor. in pressurized water and boiling water atomic reactors operating at about 550°F for a number of years. As the life of such equipment is extended or operating temperatures are raised, the possibility of embrit-tlement becomes of increasing concern to materials engineers. Much investigation work was done with respect to the use of 17-4 PH at 550°F. K. C. Antony' states "Such estimation" (of an activation energy for the diffusion of chromium in iron) "would indicate W. C. CLARKE, Jr. is Senior Research Engineer, Advanced Materials Division, Armco Steel Corp., Baltimore. Md. This manuscript is based on a talk presented at the symposium on New Developments in Stainless Steel, sponsored by the IMD Corrosion Resistant Metals Committee, Detroit, Mich., October 14-15, 1968. significant secondary aging is improbable at temperatures less than 700°F within normal component service life". This statement is modified however by the recognition of the accelerating tendencies of applied stress and internal stress as well as the possible effect of irradiation. In this investigation of 17-4 PH, the H 1100 (1900°F-1 hr oil quench or air cool + 1100°F-4 hr-air cool) condition was used, partly because this condition is normally used in atomic reactors. As shown later, the precipitation hardening temperature has no real bearing on the rate of embrittlement. An exposure temperature of 800°F was selected since embrittlement at 800°F is rapid, permitting development of relative data in time periods of 400 to 500 hr. For those not familiar, a nominal present day analysis of 17-4 PH is: C Mn P S SiCrNiCuCbN 0.04 0.30 0.015 0.015 0.60 16.00 4.30 3.25 0.23 0.030 TYPICAL BEHAVIOR OF 17-4 PH Figs. 1 and 2 show the behavior of a commercial heat of 17-4 PH under the conditions defined. Characteristically, 800°F exposure causes a rapid drop in Charpy V-notch impact strength. Tensile and yield strengths gradually increase and a gradual loss of elongation and reduction of area occur, accompanied by an increase in hardness. Notched tensile strength increases to 125 hr exposure and then sharply decreases after exposure for 500 hr. The notched vs un-notched tensile ratio remains virtually constant to exposure for 125 hr (1.67 to 1.56) but drops to 1.15 after 500 hr. Tensile ductility is not alarmingly affected even after much longer exposure times than these. For example, samples aged at 1100°F for 1 hr exposed at 800°F for 4000 hr showed a drop in elonga-tion of from 13 to 11 pct and in reduction of area of from 58 to 37 pct. Notched impact is the property SmIgth KSI 3JJ[/__________^UTS-Nrteh 260 ;/- 240 —^ 200 UTS-Smooth________._, o ioo mo Sob" m Too Hours Exposure Fig. 1—Effect of exposure at 800°F for various times on notched tensile strength and smooth tensile and 0.2 pct yield strength of 17-4 PH in the H 1100 condition.

Jan 1, 1970

By Eugene A. Thiers, William V. Morris

INTRODUCTION Direct reduction processes reduce the various commercial forms of iron oxide (pellets, concentrate, fines, etc.) to metallic iron at temperatures lower than that of molten iron. Thus, this technology includes practically, all iron reduction processes other than blast furnaces and electric pig iron furnaces (whose output in terms of world production is negligible). The product of these processes which is known as direct reduced iron (DRI), or sponge iron, is primarily used as a source of metallic iron in steel-making operations. Interest in DRI, which has been significant since the early 1960s, increased significantly in recent years with the rapid growth of DRI installed capacity throughout the world. The importance of the subject for the Circum-Pacific region stems directly from the influence that DRI has on iron ore consumption and on future steel development for this region. Although there is widespread agreement that electric furnaces will continue to increase their share of global steel output, and especially so in the countries of the Pacific Steel community, some doubts exist about future scrap supplies being adequate to support growth at past rates. The authors believe that such doubts are soundly based. As this paper points out, the total supply of all metallics used in electric furnaces may not be adequate to support the extrapolated rapid growth in electric furnace steel production. This paper seeks to provide perspective on the global and Circum-Pacific prospects for DRI in light of recent energy price developments and the current recession. In this regard, the demand for DRI within the context of recent evolutionary patterns in steel-making, the outlook of DRI supply in terms of prevailing production costs, and the prospects of new technology are discussed. THE DEMAND FOR DRI Although several reports published in the last 10 years predict high rates of growth in DRI, the subject remains a controversial one. Significant growth has indeed occurred, but not to the extent anticipated in the studies summarized in Table 1. The substantial difference between previous expectations and present reality can be ascribed primarily to: (1) lower growth in steel production than formerly anticipated; (2) numerous cancellations of DRI facilities that were previously announced; and (3) a fundamental and probably irreversible change in the economics of DRI production. Note that DRI capacity at the end of 1980 was about 16 million tonnes, a significantly lower figure than any of the projections above. In addition, DRI production was only about half of capacity, reflecting the abnormally low rates of capacity utilization in this industry. [ ] Before examining the current outlook in steel, it is pertinent to note that the market for DRI is usually different in the industrialized countries of the West from that in developing countries. In the former, the available infrastructure and industry's diversification extends DRI's potential markets to numerous steelmaking, foundry, and other industrial applications, although competition from scrap and other forms of metallic iron is constant. Scrap is generally available in these countries and, therefore, DRI competes with it in electric furnace steelmaking, basic oxygen steelmaking (as a coolant), cupola foundry operations, or as an additive in the metallic charge for open hearth and blast furnaces. On the other hand, DRI in developing countries is often allocated exclusively to domestic electric furnace steelmaking or, when capacity exceeds domestic captive requirements, to export. Notwithstanding quality considerations, DRI is being and is likely to continue to be used predominantly as a source of metallics in iron and steel-making. Other uses of DRI, such as in copper cementation represent a marginal market in terms of overall tonnage and can be ignored at this point. Therefore, DRI demand is-determined by the overall availability of metallic scrap in its various forms--a function of steel production and its probable evolution. The Global Steel Outlook Given the present recession, an objective appraisal of the long-term outlook for steel is particularly difficult. On the one hand, historical trends and, especially, the inertial forces associated with such a basic industry as steel must be recognized; such trends suggest that the current stagnation with

Jan 1, 1982

By D. K. Das, P. A. Beck, S. P. Rideout

IN a previous publication1 1200°C isothermal phase diagram sections were given for the Cr-CO-Ni, Cr-Co-Fe, Cr-Co-Mo, and Cr-Ni-Mo ternary systems, in which the a phase formed narrow, elongated solid solution fields. The present investigation is concerned with the 1200°C isothermal sections of the Co-Ni-Mo, Co-Fe-Mo, and Ni-Fe-Mo ternary systems. A prominent feature of these systems is the presence of narrow, elongated µ phase fields. The crystal structure of the phase designated as µ both here and in the previous publication1 was determined by Arnfelt and Westgren.2 For the (CO, W)µ phase, named by them Co,W, (and also frequently designated as a), these authors found that the crystal system is hexagonal-rhombohedra1 and the space group is D53d — R3,. Westgren and Mag-neli3 later found that isomorphous phases exist in the Fe-W and the Fe-Mo systems (these phases are often referred to as < and E, respectively). Henglein and Kohsok4 stated that the phase described by them as Co7Mo,; (otherwise frequently designated as c) is also isomorphous with the above three. The Co-Fe-Mo system was investigated at 1300°C by Koester and Tonn,5 who found a continuous series of solid solutions between (Co, MO)µ and (Fe, MO)µ Koester6 also indicated similar uninterrupted solid solutions in the Ni-Fe-Mo system. However, since the Ni-Mo binary system does not have a phase isomorphous with F, Koester&apos;s diagram is expected to be erroneous. No data appear to be available in the literature concerning the Co-Ni-Mo system. The face-centered cubic (austenitic) solid solut,ions of iron, nickel, and cobalt, which are quite extensive in all three systems at 1200°C, are here designated as the a phase. The body-centered cubic (ferritic) solid solutions, based on iron, are designated in this report as the ? phase, in conformity with the nomenclature used previously.&apos; Experimental Procedure The alloys were prepared by vacuum induction melting in zirconia and alumina crucibles. The lot analyses for the metals used have been given.&apos; The number of alloys prepared was 46 for the Co-Ni-Mo system, 65 for the Co-Fe-Mo system, and 113 for the Ni-Fe-Mo system. The compositions of these alloys were selected with due regard to maximum usefulness in locating phase boundaries. The alloy specimens were annealed at 1200°C in an atmosphere of purified 92 pct helium and 8 pct hydrogen mixture. Alloys consisting almost entirely of the face-centered cubic austenitic a phase, or of the body-centered cubic ferritic c phase were double-forged with intermediate annealing. The double-forged specimens were then final annealed for 90 hr at 1200 °C and quenched in cold water. Alloys containing considerable amounts of any of the other phases could not be forged. Such specimens were annealed for 150 hr at 1200°C and quenched. Microscopic specimens of all alloys were prepared by mechanical polishing, in many cases followed by electrolytic polishing. Description of the polishing and etching procedures used and tabulation of the intended compositions of the alloys prepared are being published in two N.A.C.A. Technical Notes.7,8 , Many of the alloys were analyzed chemically and, in general, the results are in excellent agreement with the intended compositions. X-ray diffraction samples were prepared by filing or crushing homogenized alloy specimens and by reannealing the obtained powders in evacuated and sealed quartz tubes. After annealing for 30 min at 1200°C the tubes were quenched into cold water. X-ray diffraction patterns were made with unfiltered chromium radiation at 30 kv, using an asymmetrical focusing camera of high dispersion. X-ray diffraction and microscopic methods were used jointly to identify the phases present in each specimen. The amounts of the phases in each alloy were estimated microscopically. The phase boundaries were located by the disappearing phase method. The results were used to construct 1200°C isothermal sections for the three ternary phase diagrams. The accuracy of the location of the phase boundaries determined in this manner is estimated to approximately ±1 pct of each component. The portion of the three phase diagrams lying between the µ, P, and 6 phases on the one hand, and the molybdenum corner on the other, has not been investigated. Recently Metcalfe reported0 a high temperature allotropic form of cobalt on the basis of dilatometric results and of cooling curves. In the present work no attempt was made to search for the new phase in the cobalt corner of the Co-Fe-Mo and Co-Ni-Mo systems. No alloy was prepared with more than 80 pct Co; the alloys used were intended to locate the boundary of the a phase saturated with cL. The microstructures of the quenched a alloys near the cobalt COrner gave no suggestion of an in-suppressible transformation On quenching. The location of the boundaries of the a + ? two-phase fields in the Fe-Ni-Mo and Fe-CO-MO systems was determined entirely by the microscopic method. The face-centered cubic a alloys near the ? field transform partially or wholly into the body-centered cubic ? phase on quenching from 1200°C to room temperature. The ? formed in this manner has an

Jan 1, 1953

By P. L. De Bruyn, R. J. Charles

THE transfer of kinetic energy of translation into other forms of energy by impact is a fundamental process in most crushing and grinding operations. During and after the impact process the original source energy may be accounted for in any of the following possible forms: 1) Kinetic energy of translation of both the impacted and impacting objects. 2) Kinetic energy of vibration of the components of the impact system. 3) Potential energy as strain energy of the components of the system or in the form of residual stresses. 4) Heat generated by internal friction during plastic deformation or during damping of elastic waves. 5) New surface energy of fractured materials. At any instant during the impact process only the strain energy of the components of the system can contribute directly to the brittle fracture process. If fracture is the desired result, as in comminution, it would seem advantageous to choose or arrange the conditions of impact so that a maximum amount of the original kinetic energy could be converted to strain energy at some moment during a single impact. The present work deals with determination of these desirable conditions for a simple case of impact and application of the principles involved to general cases of impact. Experimental Method: Longitudinal impact of a rod with a fixed end was chosen as the impact system for investigation. The rod was mounted horizontally and the fixed end was formed by butting one end of the rod against a rigidly mounted steel anvil. The rod, of pyrex glass, was 10 in. long by 1 in. diam with both ends rounded to a 6 in. radius. The rounded ends permitted reproducible impacts on the free end of the rod and assured a symmetrical fixed end. Pyrex was selected as the rod material because of the marked elastic properties of such glass and the similarity of fracture between pyrex and many materials encountered in crushing and grinding operations. The frequency of natural longitudinal oscillation of the rod was 10 kc, and thus simple electronic equipment could be used for observation of strain changes occurring in the rod at this frequency. As shown in Fig. 1, impacts on the free end of the rod were obtained either by a pendulum device or by a spring-loaded gun. Relatively heavy hammers (100 to 600 g) of mild steel were used in the pendu- lum impacts, while fairly light projectiles (20 to 80 g) were fired from the spring-loaded gun. One of the main objects of the experimental work was to obtain the strain-time history of the rod as a function of the mass and kinetic energy of the impacting hammers. For this purpose a technique involving wire resistance strain gages and a recording oscilloscope was employed. Five gages were applied at equidistant sections along the rod, and by means of a switching arrangement the strain-time history at any section, and for any impact, could be obtained in the form of an oscillograph with a time base. The equation relating strain and voltage change across a strain gage through which a constant current is flowing is as follows: e = ?v/iRF [1] ? = strain, ?v = voltage change, i = gage current, R = gage resistance, and F = gage factor (from manufacturer&apos;s data — SRA type, Baldwin Lima Corp.). With the above equation an oscillograph depicting voltage change vs time on a single trace can be converted directly to a strain-time diagram if a calibration of the vertical response on the oscilloscope screen for specific voltage inputs is available. In the present case the calibration was obtained by photographing precisely known audio frequency voltages on the same oscillograph as that on which a voltage-time trace from a strain gage had been made. Synchronization of the beginning of the single trace with the beginning of the impact was accomplished by permitting contact of the impacting objects to close an electrical circuit from which a voltage pulse, sufficient to initiate the trace, was obtained. The struck end of the rod was lightly silvered for purposes of electrical conduction so that it would form one of the electrical contacts. Markers every 100 micro-seconds on the traces served for a time base calibration. Determinations of the kinetic energies of translation prior to impact were made in the case of the pendulum hammers by measuring the height of fall of the hammer and in the case of the projectiles by measuring the exit velocity from the gun barrel by means of an electrical circuit employing light sources, slits, and phototubes.&apos; During the experimental work it became evident that the time of contact between the impacting object and the rod was an important variable in the impact process. Measurements of the times of contact were made, therefore, for every impact for which a strain-time record was obtained. The time of contact was determined by permitting the impacting components, when in contact, to act as a closed switch and discharge a condenser at relatively constant voltage. The discharge was observed and photographed with a time base on the oscilloscope screen.

Jan 1, 1957

By C. T. Butler, W. W. Kay, R. D. Boddorff, R. L Ash

DRILLING and blasting in anthracite open-pit mines is a continuous problem to contractors and explosive engineers because of the diverse conditions caused by the nature of the geological formations, the extensive mining of the portions of coal beds near the surface, and the proximity of many strip pits to populated areas. Pennsylvania anthracite occurs in four separate long and narrow fields totaling only 480 sq miles. The coal measures are rock strata and coal beds that are considerably folded and faulted. The crests of the anticlines are eroded extensively. The beds outcrop on the mountain sides and dip under the valleys. At first only the upper portions of the syn-clines could be stripped. Now stripping to increasingly greater depths is economically possible, as is indicated by the fact that the proportion of freshly mined anthracite produced by strip mining has increased from 3.7 pct of the total tonnage in 1930 to 29.6 pct in 1950. Much of the rock overlying the deeper beds now being stripped is so extensively broken that considerable difficulty is experienced in drilling satisfactory blast holes and in using explosives in such manner as to insure a uniformly broken material easily removed by the excavating machinery. Such breaking of rock strata has occurred because the bed now being stripped has been mined extensively in former years by underground methods, and tops of gangways and chambers have subsequently failed. Draglines are used to uncover coal where the overburden can be moved with little or no re-handling. These machines range in size from those having a 2 cu yd capacity bucket on a 60-ft boom to those handling a 25 cu yd bucket on a 200-ft boom. Draglines are also used to strip to the bottom of the coal basins if the depth and the distance between the crops are not too great. For this type of operation blast holes are drilled full depth to the bed. These holes are commonly 30 to 90 ft deep; however, in exceptional cases, holes may be as shallow as 12 ft or as deep as 130 ft. Drilling is normally done for blasts of 12,000 to 60,000 cu yd of overburden, 30,000 cu yd being considered an average blast if vibration is not the controlling factor. Where the stripping of wide basins or the exposure of a moderately pitching vein makes the use of draglines impractical, dipper front shovels equipped with 4 to 6 cu yd buckets load into trucks. Overburden is removed in benches of 25 to 30 ft with blast holes drilled 4 or 5 ft deeper than the planned floor of the bench. For shovels under 5 cu yd bucket capacity the volume blasted varies from 8000 to 12,000 cu yd, whereas a volume of 30,000 to 50,000 cu yd of overburden is frequently blasted at one time for the larger shovels where vibration is not an important factor. During the past decade the churn drill, generally the Model 42-T Bucyrus-Erie blast hole drill equipped for drilling 9-in. diam holes, has become the most common blast hole drilling machine. Electricity powers half the churn drills in use and is preferred on the large strippings where electric shovels are operated and the working area is concentrated. On these operations the cost of additional electricity for the drills is less than the cost of fuel to operate diesel units because of the existing large demand load of the excavating equipment. Moreover, electric motors start more easily in cold weather and generally are less expensive to maintain. Diesel driven units are employed where a higher degree of mobility is required. The average drilling speed is 8 ft per hr, although in softer rocks a rate of 15 ft per hr is attained. Where rock is hard and strata is badly broken, drill speeds may be less than 2 ft per hr. Low drilling production results under these circumstances when loose material falling from the upper portion of the drill holes causes drill stems to be jammed. Rock formations vary so greatly in the region that a 9-in. diam churn drill bit may become dull after drilling only 2 ft or may drill satisfactorily for 56 ft; however, an average of 35 ft is usual in sandstone of medium hardness. Dull bits are hoisted to flat bed trucks by the sand line of the drill and are usually sharpened in the contractor&apos;s bit shop adjacent to the job. Care is generally taken to cover the thread end of the bit with a cap. To facilitate handling of bits around the drill, a heavy thread protector having an eye top is becoming more popular than the flat-top rubber or metal cap furnished with new bits. The 9-in. diam blast holes for a 25 to 30 ft bench are normally on 18x18 ft to 20x20 ft spacings, depending on the character of the overburden, although in broken ground 15x18 ft centers may be used to obtain better breakage and a more even bottom for the bench. The patterns of holes for shots

Jan 1, 1953