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By Peter R. Drummond

THE final casting of refined copper has been re-J- stricted for generations by the following sequence of operations: Filling the reverberatory furnace, melting, skimming, blowing or flapping, and poling. The hoped-for 24 hr cycle, producing 300 tons or more, has been taken up largely with the necessary bat time-consuming tasks of cleaning the bath, sulphur elimination, and in turn removal of excess oxygen to produce tough-pitch copper. Incidental to comparatively slow melting under combustion gases, copper oxides react with the furnace lining, and the slag so-formed must be completely recycled. The three-phase arc furnace has eliminated some of the cycle stages, and telescoped the remainder into a continuous operation. Electrical energy, supplied to graphite electrodes enclosed in high grade refractories, rapidly melts copper cathodes and sustains a stream of metal, containing approximately 0.01 pct oxygen, without contamination from fuel. The arc was struck on the first large electric furnace for melting copper in the United States on April 13, 1949. The earliest use of this type of furnace was at Copper Cliff, Ont., in 1936, and an admirable description of their installation has been published? Copper, melted in the Baltimore furnace, is used to cast billets, and the installation differs somewhat from the Canadian, as will be described. The arc furnace is a heavy-duty, three-phase furnace, holding 50 tons, the general outline of which appears on Fig. 1. The steel shell is 15 ft ID with a bottom radius of 14 ft 2 in. The roof, separate and distinct from the body, consists of a 15-ft water-cooled, cast-steel ring of the same outside diameter as the furnace. The center line of the furnace lies 9 ft 6 in. from that of the trunnions, permitting a 5" backward tilt for skimming, and a 40" maximum nose tilt forward for complete draining. Normally, the furnace overflows by displacement, and the use of the forward tilt arrangement is restricted to covering charging delays. The charging slot, 3 ft 8 in. x 5 in., lies on the north center line, the tap hole on the south, and the 30x30 in. skim door 45" to the west of the slot. The original 20-in. graphite electrodes were replaced with 14 in. in December 1949. Three conventional direct current winch drives, governed by electrical controls, position each electrode which has individual mast supports and counterweights. An independent circulation supplies cooling water for the electrode glands, the roof ring, charge slot, and the skim door frame. Arc Furnace Refractories Hearth: Fused-in monolithic bottoms had been used in copper arc furnaces, installed prior to April 1949. These consisted of thin layers of periclase, successively fused in place over preliminary brick courses. Heat was obtained from the arc, using a T-like arrangement of broken electrodes resting directly on the periclase to be fused. The operation, taking weeks to perform, was very expensive. The chemically-bonded magnesite-brick bottom, installed at Baltimore, was the first of its kind and a radical departure from previous practice. It consists of a 1 to 6-in. layer of castable refractory laid on the steel shell, modifying it to a 12 ft 2 in. bottom radius. Two courses of 9x2 % -in. fireclay straights and keys follow. The third course is made of 9-in. magnesite blocks of special shape to form circles of an inverted arch. It was started by a four piece keystone with skew-backs forming the outer course. The fourth course also started on a central keystone, or button, of four 90" segments, 12 in. diam x 13 Vz in. deep, and continued with 13%-in. blocks. Skewbacks at the shell completed the course to produce a horizontal surface for the side walls with a single course of No. 2 arch fireclay against the steel. Dry chrome-magnesite cement was brushed over each course after laying, and a 1-in. expansion space between the brick and the shell was filled with the same mixture. The total bottom thickness, excluding the castable material, was 5 in. of clay plus 22% in. of chemically-bonded magnesite. Tap Hole: A 5-in. OD and 3-in. ID silicon-carbide tube constitutes the tap hole and is set tangential to the upper course of the furnace bottom. Molten metal fills the tube when the furnace is level and filled to capacity. Side Walls: The lining, against the shell, consists of a 9x4Y2x3 in. soldier course of fireclay, using straights and No. 1 arches to turn the circles. A second soldier course of 9x4'/2x2'/2-in. fireclay was laid in a somewhat similar fashion. Three courses of 13Y2x6x3 in. and 9x6~3 in. of final magnesite, laid flat, completed the lining, using Nos. 1 and 2 keys to turn the circles. Cardboard spacers were placed between every two bricks in horizontal courses, and a thin coat of chrome-magnesite cement filled the joints between the firebrick and magnesite. A sprung-arch spanned the skim door with jambs of suitable magnesite shapes. Charge Slot: The slot is 3 ft 8 in. wide x 5 in. high. A silicon-carbide sill of special shapes has a 30" slope to allow cathodes to slide easily into the bath. The original arch was flat, and composed of Nos. 1 and 2 wedge magnesite with a 6-ft radius. It projected 5 in. over the sill, and, being a flat arch, gave an 18 15/16-in. opening between the inner edge and the metal line. The whole assembly was later raised 9 in., and the flat arch replaced with an arch, the lower edge of which maintained the 5-in, width from the outer to inner edges as shown in Fig. 2. A water-cooled, cast-copper jacket protects the steel shell behind the slot.

Jan 1, 1952

By E. H. Crabtree

THE heavy-media separation plant at the Central mill of the Eagle-Picher Mining and Smelting Co., near Picher, Okla., was started in February 1939. Since that time twenty-four million tons of lead-zinc ore has been treated in the mill, of which approximately 70 pct has been treated by heavy-media separation. Until January 1945, galena concentrate from the flotation plant was used as the sink-float medium. At that time minus 100-mesh ferrosilicon was substituted for the galena, and its use has been continued. The purpose of this paper is to compare the operating results and the costs of operation of the two media. For the purpose of this comparison the period Jan. 1, 1943, to Jan. 29, 1945, has been taken to represent the operation using galena medium; the period Jan. 29, 1945, to Sept. 1, 1946, to represent the ferrosilicon medium. During the first of 'these periods about seven and one-half million tons was milled; during the latter, or ferrosilicon period, more than five million tons was milled. It is believed that these tonnages are large enough to give representative data. FLOWSHEET The Central mill has a capacity of 15,000 tons per day. Briefly, the treatment used consists of crushing the ore through 11/2-in. and wet-screening out the minus 3/16in. portion. The minus 1 1/2 plus 3/16in. washed product is treated by the heavy medium; the minus 3/16-in. product is deslimed and jigged. The heavy-media cone tailing is discarded and the cone .concentrate is recrushed and jigged for the production of a coarse lead concentrate. The cone concentrate, impoverished of lead, is then ground in ball mills and treated by differential flotation. The minus 3/16-in. jig product is similarly treated. HEAVY-MEDIA PLANT The heavy-media cone plant treating the minus 1 ½ plus 3/16-in. product consists of one open-top cone 20 ft in diameter by 20 ft deep, equipped with a central airlift. Concentrate is discharged from the top of the airlift to a 3 by 14-ft low-head Allis-Chalmers drainage screen. Medium drains back into the cone, from the first half of this screen. The last half of the screen is provided with washing sprays for further removal of medium. Cone tailing is similarly drained and washed. Undersize from the washing screen is thickened and then cleaned in the medium-cleaning plant. When galena was used as a medium, cleaning was done in Fagergren and Denver flotation machines. The ferrosilicon is cleaned in three 36-in. and one 24-in. Crockett magnetic separators. A flowsheet of the cone plant is shown in Fig 1.

Jan 1, 1947

By William L. Russell

A new instrument for measuring bulk and grain vol-urnes is described, and compared in accuracy with two modifications of Nutting's method, All three methods are reasonably fast, and in all three the probable error in the determinations of bulk volume can be kept down to about 0.1 to 0.2 per cent. INTRODUCTION The new pycnometer described in this paper is designed to give speed without sacrificing accuracy. Some time ago Russell1 described a somewhat similar appa-ratus; however, the instrument described in the present paper is more accurate. Nutting' designed the pycnom-eter shown in Fig. 1 (A and B), and Pirson3 has shown points like those of Fig. 1-D. The procedure described in this paper for using Nutting's pycnometer has been modified from that used by Nutting. CHOICE OF LIQUIDS The best liquids for measuring grain volumes are hydrocarbons or chlorinated hydrocarbons which evaporate slowly. Trimethylbenzene or mesitylene, triethylben-zene, carbon tetrachloride, and tertalin have been used for this purpose. Any of these liquids or mercury may be used for measuring bulk volume. PREVENTING ERRORS DUE TO INVASION OF LIQUID Errors due to the invasion of the porous, permeable bulk samples by the liquid of measurement may be prevented or greatly reduced by using mercury, by applying an impervious coating, or by soaking the bulk samples in the liquid of measurement until bubbles cease issuing from them. Errors are produced only by the invasion of the samples which occurs between movement when the samples are weighed in air and the moment the weight of liquid displaced by the sample in the pycnometer is determined. No error is produced by varying degrees of filling of the pore space of the samples by the liquid of measurement. The excess liquid should be wiped from the samples immediately before the weight in air W1, is determined, and they should be placed in the pycnometers immediately. OPERATION OF NEW INSTRUMENT The new pycnometer is shown in Fig. 2. It consists of two glass parts which fit together at a ground glass joint. The spaces J and D are connected by fine glass tubes E, and the space J must be larger than the space D by at least the volume of the largest sample to be tested. The projections H afford a grip for twisting when the two parts tend to stick together. The position of the sample in mercury is shown by I, and in other liquids by M. F is a permanent mark on the glass, and G is a tem-Some marker. In order to determine the position at which G should be placed, invert the instrument so that A is at the bottom, and fill space D. Place the cub K

Jan 1, 1958

By E. H. Crabtree

The heavy-media separation plant at the Central mill of the Eagle-Picher Mining and Smelting Co., near Picher, Okla., was started in February 1939. Since that time twenty-four million tons of lead-zinc ore has been treated in the mill, of which approximately 70 pct has been treated by heavy-media separation. Until January 1945, galena concentrate from the flotation plant was used as the sink-float medium. At that time minus 100-mesh ferrosilicon was substituted for the galena, and its use has been continued. The purpose of this paper is to compare the operating results and the costs of operation of the two media. For the purpose of this comparison the period Jan. I, 1943, to Jan. 29, 1945, has been taken to represent the operation using galena medium; the period Jan. 29, 1945, to Sept. I, 1946, to represent the ferrosilicon medium. During the first of these periods about seven and one-half million tons was milled; during the latter, or ferrosilicon period, more than five million tons was milled. It is believed that these tonnages are large enough to give representative data. Flowsheet The Central mill has a capacity of 15,000 tons per day. Briefly, the treatment used consists of crushing the ore through I½- in. and wet-screening out the minus 3/16-in. portion. The minus 136 plus 3/16-in. washed product is treated by the heavy medium; the minus 3/16-in. product is deslimed and jigged. The heavy-media cone tailing is discarded and the cone concentrate is recrushed and jigged for the production of a coarse lead concentrate. The cone concentrate, impoverished of lead, is then ground in ball mills and treated by differential flotation. The minus 3/16-in. jig product is similarly treated. Heavy-media Plant The heavy-media cone plant treating the minus I½ plus 3/16-in. product consists of one open-top cone 20 ft in diameter by 20 ft deep, equipped with a central airlift. Concentrate is discharged from the top of the airlift to a 3 by 14-ft low-head Allis-Chalmers drainage screen. Medium drains back into the cone, from the first half of this screen. The last half of the screen is provided with washing sprays for further removal of medium. Cone tailing is similarly drained and washed. Undersize from the washing screen is thickened and then cleaned in the medium-cleaning plant. When galena was used as a medium, cleaning was done in Fagergren and Denver flotation machines. The ferrosilicon is cleaned in three 36-in. and one 24-in. Crockett magnetic separators. A flowsheet of the cone plant is shown in Fig I.

Jan 1, 1949

By Walter Miller

Petroleum refining, like other industries in the United States in 1940, focused much attention on its duties and opportunities in the field of national defense. In crdiiter-distinction to the situation during the World War, the industry is well constituted as to existing equipment, processes, and capacity, together with planned additions, to handle any demands that may be made upon it by the defense program needs. Evident throughout was the growing sense of ability to manipulate oil molecules for the production of materials not hitherto considered as being available commercially from petroleum. The development and application of catalytically induced reactions on a broad front is one of the distinguishing characteristics of refining progress during the year. In the hands of the refiner, catalysis has developed into a most important means to secure higher antiknock motor fuels and 100-octane fighting grade aviation gasoline in greatly increased quantity. One illustration of the importance being assigned to catalytic refining is the construction by the Houdry group of an experimental laboratory at a cost of $7'307'30'3, to be staffed with over 100 research men7 solely for the purpose of studying catalytic reactions. In the aviation gasoline picture a modification of specifications by the Government widened the field from which acceptable fuel can be secured, and the continued expansion of the alkylation process in itself promises to go far in filling the demands expected from 50,000 fighting planes. It is conservatively estimated that eyen now the industry can turn out 1,300,000 gal. of 100-octane aviation gasoline daily. The upper limits of octane numbers for regular grades of motor gasoline reached in 1939 were not exceeded, but the average octane number approached the upper level more closely. A survey made of the 1940-model cars showed no rise in average octane requirement of the engines. The trend throughout the year was for greater reliance on the A.S.T.M. method for octane determination. The "1939 Research Method" is less widely used, although still considerably in evidence; the older L-3 method seems to have lost considerable ground and to be on the way out. Re-forming Capacities Increased Interest continued to grow apace in re-forming low-octane naphthas into high-octane gasolines. Some additions, although not great, were announced in thermal re-forming capacity. Interest in catalytic re-forming was greatly stimulated by the technical success of the operation of the converted hydrogenation unit at the Bay-way, N. J., rehery of the Standard Oil Company of New Jersey, mentioned in last year's review, and also by the knowledge that recoverable proportions of toluene are present in the hydroformed (catalytically

Jan 1, 1941

By Walter Crafts, C. M. Offenhauer

In a previous study of the carbide phase of chromium steels, it was shown that chromium carbide (Cr7C3) is 2 more stable carbide than cementite (Fe3C) at tempering temperatures above about 500°C. in quenched steels containing from about I to 7.5 per cent chromium. Below this temperature the carbide phase was found to consist of cementite in all of the steels examined. In the present investigation it was found that a similar change in the carbide phase takes place in molybdenum and in chromium-molybdenum steels at approximately the same tempering temperatures. The approximate ranges of occurrence of the alloy carbides Cr7C3, Cr4C, and Mo2C have been determined in steels containing up to 5 per cent chromium, I per cent molybdenum and 0.60 per cent carbon. The experimental steels were made in small high-frequency furnaces from Armco ingot iron, low-carbon ferrochromium and calcium molybdate and were poured into 2-in. ingots, which were forged to I-in. round bars. Sections of the bars were heated to the carbide-solution temperature and cooled rapidly to induce austenite transformation at relatively low temperatures. The steels were subject to some segregation, therefore comparatively high temperatures were necessary to dissolve all of the carbides. The microstructure of the quenched specimens was examined to verify the belief that the carbides had been dissolved, and also to determine the nature of the nonmartensitic structure of incompletely hardened steels. Ferrite and pseudo-martensite were present in some of the specimens, but all specimens were quenched at such a rate that pearlite was not formed, in order to ensure the formation of the carbides at temperatures lower than about 500°C. It had been found previously1 that in chromium steels the chromium carbide could be produced by transformation of austenite in the pearlite range above 500°C., and, in order to start the tempering treatment with the carbide present as cementite, the absence of pearlite was considered to be essential. The specimens were tempered for 64 hr. to permit the formation of the stable carbide. After tempering, the specimens were submitted to microscopic and X-ray examination. The carbides were separated either electrolytically in a 5 per cent HCl electrolyte or with an acidified copper-potassium chloride solution. The crystal structure of the carbide phases was determined by the Debye-Scherrer type of X-ray patterns. The X-ray patterns obtained from specimens near the boundaries between two types of carbides were generally weak and diffuse as compared with the other patterns. For this reason, and because the more soluble components may be masked, the boundaries between the phases could not be determined precisely. However, the X-ray analysis of electrolytically separated carbides has given consistently reproducible results

Jan 1, 1943

By Walter Miller

Petroleum refining, like other industries in the United States in 1940, focused much attention on its duties and opportunities in the field of national defense. In crdiiter-distinction to the situation during the World War, the industry is well constituted as to existing equipment, processes, and capacity, together with planned additions, to handle any demands that may be made upon it by the defense program needs. Evident throughout was the growing sense of ability to manipulate oil molecules for the production of materials not hitherto considered as being available commercially from petroleum. The development and application of catalytically induced reactions on a broad front is one of the distinguishing characteristics of refining progress during the year. In the hands of the refiner, catalysis has developed into a most important means to secure higher antiknock motor fuels and 100-octane fighting grade aviation gasoline in greatly increased quantity. One illustration of the importance being assigned to catalytic refining is the construction by the Houdry group of an experimental laboratory at a cost of $7'307'30'3, to be staffed with over 100 research men7 solely for the purpose of studying catalytic reactions. In the aviation gasoline picture a modification of specifications by the Government widened the field from which acceptable fuel can be secured, and the continued expansion of the alkylation process in itself promises to go far in filling the demands expected from 50,000 fighting planes. It is conservatively estimated that eyen now the industry can turn out 1,300,000 gal. of 100-octane aviation gasoline daily. The upper limits of octane numbers for regular grades of motor gasoline reached in 1939 were not exceeded, but the average octane number approached the upper level more closely. A survey made of the 1940-model cars showed no rise in average octane requirement of the engines. The trend throughout the year was for greater reliance on the A.S.T.M. method for octane determination. The "1939 Research Method" is less widely used, although still considerably in evidence; the older L-3 method seems to have lost considerable ground and to be on the way out. Re-forming Capacities Increased Interest continued to grow apace in re-forming low-octane naphthas into high-octane gasolines. Some additions, although not great, were announced in thermal re-forming capacity. Interest in catalytic re-forming was greatly stimulated by the technical success of the operation of the converted hydrogenation unit at the Bay-way, N. J., rehery of the Standard Oil Company of New Jersey, mentioned in last year's review, and also by the knowledge that recoverable proportions of toluene are present in the hydroformed (catalytically

Jan 1, 1941

By Edward Saibel, W. T. Lankford

DURING the course of an investigation of the plastic flow of aluminum aircraft sheet under combined loads, several problems arose in which analyses of the conditions leading to unstable plastic flow were needed. In order to provide these analyses, and in the hope of attaining a better understanding of the general problem of unstable plastic flow, the present study was undertaken. It is well known that unstable plastic flow is associated with a form of mechanical instability frequently observed during the plastic deformation of ductile metals. This instability is characterized by the fact that, at its onset, plastic straining will proceed at a constant magnitude of the externally applied forces. It is further typical that the resulting unstable flow is generally accompanied by the formation of a neck, a local bulging, or some other form of heterogeneous deformation. Perhaps the most familiar example of unstable plastic flow to the metallurgist is the formation of the neck in the ordinary tension test. It is well known that in this test most metals exhibit a maximum in the externally applied load, followed by the process commonly referred to. as "necking down." Instability in the simple tension test has been frequently discussed and numerous investigators 1,2.3,12 have derived the criterion for instability in terms of a critical rate of strain hardening. It is the purpose of the present study to extend the criterion for unstable flow to the more general case where the body undergoing flow is subjected to combined loads. It is hoped that the problems considered will provide a more consistent picture of the general features of this important form of mechanical instability. The most practical implications of instability in plastic flow arise in connection with sheet metal forming operations. In the formation of a great many sheet metal parts, the deformations are produced primarily by stretching. If the conditions which lead to unstable flow are fulfilled in such a formation, and if the resultant instability leads to a pronounced localization of flow, then the useful limit of ductility of the material will have been reached. It is also essential that the effects of instability be taken into consideration in any study of the plastic flow and rupture of metals under combined stresses. In general, when unstable plastic flow is initiated, straining becomes localized, the geometry of the test specimen changes, and the state of stress in that portion of the specimen undergoing plastic flow is no longer under external control. In many past studies of the effects of combined stresses on flow and rupture, account has not been taken of the fact that, although a given test may be started under constant

Jan 1, 1947

By J. G. Thompson

Determinations of the hydrogen content of irons and steels invariably are subject to two serious difficulties: (I) the determination of amounts of 0.001 per cent or less of any constituent requires an analytical procedure of unusual precision and accuracy, and (2) the hydrogen content of many samples is dccidedly not stable; i.e., there is no assurance that the content at the time the analysis is made is the same as it was when the sample was taken or at some previous time in the history of the piece. In respect to both the precision and speed of the determination, the vacuum-extraction procedure for the determination of hydrogen possesses advantages over the vacuum-fusion procedure, which has been the recognized standard for the detcrmination of other gaseous elements in sted-— oxygen and nitrogen for example. The basic facts upon which the vacuum-extraction procedure is based—i.e., the evolution of gas from solid steel when it is heated in vacuum, and the fact that the evolved gas is rich in hydrogen—were observed by Graham1 many years ago. Subsequent investigations have shown that hydrogen exists in steel chiefly as interstitial solid solution (i.e., in atomic dispersion) and perhaps to some extent as hydridcs that are unstable and decompose to form atomic hydrogen. In either case practically all of the hydrogen in steel at elevated temperatures exists in the atomic condition and consequently is able to diffuse readily through the metal. In contrast to this, other gaseous elements, such as oxygen and nitrogen, normally are present in steel chiefly in the form of relatively stable compounds such as oxides or silicates and nitrates; i.e., the oxygen and nitrogen atoms are fixed in stable combinations and are not free to migrate. The only oxygen or nitrogen that can diffuse in steel is the small amount in solution (atomically dispersed) and the rate of diffusion of these atoms is much less than that of hydrogen because of their greater atomic volume and lower mobility. Consequently, it is possible to extract most or all of the hydrogen from solid steel in a reasonable time at moderately elevated temperatures, but the procedure is not applicable to the determination of oxygen or nitrogen. It is only within the last decade that these observations have been developed into an analytical procedure for the determination of hydrogen. Apparatus and Procedure The apparatus used at the National Bureau of Standards,6 for the determination of hydrogen by the vacuum-extraction procedure, is shown schematically in Fig. I. The specimen, freshly cleaned and dried,* is suspended on a fine wire within the quartz furnace tube that is connected to the reservoir system and analytical train.

Jan 1, 1945

By Walter Crafts, C. M. Offenhauer

In a previous study of the carbide phase of chromium steels, it was shown that chromium carbide (Cr7C3) is 2 more stable carbide than cementite (Fe3C) at tempering temperatures above about 500°C. in quenched steels containing from about I to 7.5 per cent chromium. Below this temperature the carbide phase was found to consist of cementite in all of the steels examined. In the present investigation it was found that a similar change in the carbide phase takes place in molybdenum and in chromium-molybdenum steels at approximately the same tempering temperatures. The approximate ranges of occurrence of the alloy carbides Cr7C3, Cr4C, and Mo2C have been determined in steels containing up to 5 per cent chromium, I per cent molybdenum and 0.60 per cent carbon. The experimental steels were made in small high-frequency furnaces from Armco ingot iron, low-carbon ferrochromium and calcium molybdate and were poured into 2-in. ingots, which were forged to I-in. round bars. Sections of the bars were heated to the carbide-solution temperature and cooled rapidly to induce austenite transformation at relatively low temperatures. The steels were subject to some segregation, therefore comparatively high temperatures were necessary to dissolve all of the carbides. The microstructure of the quenched specimens was examined to verify the belief that the carbides had been dissolved, and also to determine the nature of the nonmartensitic structure of incompletely hardened steels. Ferrite and pseudo-martensite were present in some of the specimens, but all specimens were quenched at such a rate that pearlite was not formed, in order to ensure the formation of the carbides at temperatures lower than about 500°C. It had been found previously1 that in chromium steels the chromium carbide could be produced by transformation of austenite in the pearlite range above 500°C., and, in order to start the tempering treatment with the carbide present as cementite, the absence of pearlite was considered to be essential. The specimens were tempered for 64 hr. to permit the formation of the stable carbide. After tempering, the specimens were submitted to microscopic and X-ray examination. The carbides were separated either electrolytically in a 5 per cent HCl electrolyte or with an acidified copper-potassium chloride solution. The crystal structure of the carbide phases was determined by the Debye-Scherrer type of X-ray patterns. The X-ray patterns obtained from specimens near the boundaries between two types of carbides were generally weak and diffuse as compared with the other patterns. For this reason, and because the more soluble components may be masked, the boundaries between the phases could not be determined precisely. However, the X-ray analysis of electrolytically separated carbides has given consistently reproducible results

Jan 1, 1943

By Walter Crafts, C. M. Offenhauer

IN a previous study' of the carbide phase of chromium steels, it was shown that chromium carbide (Cr7C3) is a more stable carbide than cementite (Fe3C) at tempering temperatures above about 500°C. in quenched steels containing from about r1to 7.5 per cent chromium. Below this temperature the carbide phase was found to consist of cementite in all of the steels examined. In the present investigation it was found that a similar change in the carbide phase takes place in molybdenum and in chromium-molybdenum steels at approximately the same tempering temperatures. The approximate ranges of occurrence of the alloy carbides Cr7C3, Cr4C, and Mo2C have been determined in steels containing up to 5 per cent chromium, 1 per cent molybdenum and 0.60 per cent carbon. The experimental steels were made in small high-frequency furnaces from Armco ingot iron, low-carbon ferrochromium and calcium molybdate and were poured into 2-in. ingots, which were forged to 1-in. round bars. Sections of the bars were heated to the carbide-solution temperature and cooled rapidly to induce austenite transformation at relatively low temperatures. The steels were subject to some segregation, therefore comparatively high temperatures were necessary to dissolve all of the carbides. The microstructure of the quenched specimens was examined to verify the belief that the carbides had been dissolved, and also to determine the nature of the nonmartensitic structure of incompletely hardened steels. Ferrite and pseudomartensite were present in some of the specimens, but all specimens were quenched at such a rate that pearlite was not formed, in order to ensure the formation of the carbides at temperatures lower than about 500°C. It had been found previously' that in chromium steels the chromium carbide could be produced by transformation of austenite in the pearlite range above 500°C., and, in order to start the tempering treatment with the carbide present as cementite, the absence of pearlite was considered to be essential. The specimens were tempered for 64 hr. to permit the formation of the stable carbide. After tempering, the specimens were submitted to microscopic and X-ray examination. The carbides were separated either electrolytically in a 5 per cent HCl electrolyte or with an acidified copper-potassium chloride solution. The crystal structure of the carbide phases was determined by the Debye-Scherrer type of X-ray patterns. The X-ray patterns obtained from specimens near the boundaries between two types of carbides were generally weak and diffuse as compared with the other patterns. For this reason, and because the more soluble components may be masked, the boundaries between the phases could not be determined precisely. However, the X-ray analysis of electrolytically separated carbides has given consistently reproducible results

Jan 1, 1943

By W. T. Lankford, Edward Saibel

During the course of an investigation of the plastic flow of aluminum aircraft sheet under combined loads, several problems arose in which analyses of the conditions leading to unstable plastic flow were needed. In order to provide these analyses, and in the hope of attaining a better understanding of the general problem of unstable plastic flow, the present study was undertaken. It is well known that unstable plastic flow is associated with a form of mechanical instabi!ity frequently observed during the plastic deformation of ductile metals. This instability is characterized by the fact that, at its onset, plastic straining mill proceed at a constant magnitude of the externally applied forces. It is further typical that the resulting unstable flow is generally accompanied by the formation of a neck, a local bulging, or some other form of heterogcneous deformation. Perhaps the most familiar example of unstable plastic flow to the metallurgist is the formation of the neck in the ordinary tension test. It is well known that in this test most metals exhibit a maximum in the externally applied load, followed by the process commonly referred to as "necking down." Instability in the simple tension test has been frequently discussed and numerous investigators1'2'3'12 have derived the criterion for instability in terms of a critical rate of strain hardening. It is the purpose of the present study to extend the criterion for unstable flow to the more general case where the body undergoing flow is subjected to combined loads. It is hoped that the problems considered will provide a more consistent picture of the general features of this important form of mechanical instability. The most practical implications of instability in plastic flow arise in connection with sheet metal forming operations. In the formation of a great many sheet metal parts, the deformations are produced primarily by stretching. If the conditions which lead to unstable flow are fulfilled in such a formation, and if the resultant instability leads to a pronounced localization of flow, then the useful limit of ductility of the material will have been reached. It i; also essential that the effects of instability be taken into consideration in any study of the plastic flow and rupture of metals under combined stresses. In general, when unstable plastic flow is initiated, straining becomes localized, the geometry of the test splecimen changes, and the state of stress in that portion of the specimen undergoing plastic flow is no longer under external control. In many past studies of the effects of combinetl Metals On flow and rupture) account has not been taken of the fact that, although a given test may be started under constant

Jan 1, 1947

By Edward Saibel, W. T. Lankford

During the course of an investigation of the plastic flow of aluminum aircraft sheet under combined loads, several problems arose in which analyses of the conditions leading to unstable plastic flow were needed. In order to provide these analyses, and in the hope of attaining a better understanding of the general problem of unstable plastic flow, the present study was undertaken. It is well known that unstable plastic flow is associated with a form of mechanical instabi!ity frequently observed during the plastic deformation of ductile metals. This instability is characterized by the fact that, at its onset, plastic straining mill proceed at a constant magnitude of the externally applied forces. It is further typical that the resulting unstable flow is generally accompanied by the formation of a neck, a local bulging, or some other form of heterogcneous deformation. Perhaps the most familiar example of unstable plastic flow to the metallurgist is the formation of the neck in the ordinary tension test. It is well known that in this test most metals exhibit a maximum in the externally applied load, followed by the process commonly referred to as "necking down." Instability in the simple tension test has been frequently discussed and numerous investigators1'2'3'12 have derived the criterion for instability in terms of a critical rate of strain hardening. It is the purpose of the present study to extend the criterion for unstable flow to the more general case where the body undergoing flow is subjected to combined loads. It is hoped that the problems considered will provide a more consistent picture of the general features of this important form of mechanical instability. The most practical implications of instability in plastic flow arise in connection with sheet metal forming operations. In the formation of a great many sheet metal parts, the deformations are produced primarily by stretching. If the conditions which lead to unstable flow are fulfilled in such a formation, and if the resultant instability leads to a pronounced localization of flow, then the useful limit of ductility of the material will have been reached. It i; also essential that the effects of instability be taken into consideration in any study of the plastic flow and rupture of metals under combined stresses. In general, when unstable plastic flow is initiated, straining becomes localized, the geometry of the test splecimen changes, and the state of stress in that portion of the specimen undergoing plastic flow is no longer under external control. In many past studies of the effects of combinetl Metals On flow and rupture) account has not been taken of the fact that, although a given test may be started under constant

Jan 1, 1947

By N. K. Chen, R. Maddin, K. T. Aust

Lattice rotations occurring on the tension and compression sides during the bending of molybdenum single crystals at room temperature were followed in detail. Observations were also made concerning crystallite rotations and slip traces. &apos;The results indicated the participation of (110) as slip planes and the<111>as slip directions. Crystallite fragmentation during bending was also noted. EVERAL investigations have been conducted on pure bending of face-centered cubic metals.&apos;-" Structural changes as a result of "bend-gliding," and the importance of bending and constraints during conditions of axial stressing, have been noted.&apos;," It was evident that the mechanism by which the bending deformation occurs is more complex than the case of simple shear. Studies of the bending of body-centered cubic metals from the aspect of the crystallographic mechanism appear to be lacking. Molybdenum single crystals therefore were subjected to deformation by bending in order to determine the reaction of the body-centered cubic lattice to bending stresses. It might be expected for the bending of single crystals that lattice rotation as represented by the axis of compression would proceed toward the pole of the active slip plane and that rotation of the tensile stress axis should indicate the slip direction.&apos; It was believed therefore that a study of the tension and compression sides of bent molybdenum single crystals should give data concerning the operative slip system. Specific problems which are pertinent to the overall picture of plastic deformation were also investigated, e.g., whether the constraints offered by pure bending give rise to the deformation band, and also the nature of the asterism occurring on the tension and compression sides during the stages of deformation. Experimental Procedure Eight single crystal specimens were grown from sintered molybdenum rods 143 in. diameter using the method described previously.6 The purity of this material was reported as 99.9 pct as described elsewhere.&apos; The specimens were % to 1/10 in. in diameter x 7 in. long, with single crystals approximately 1 to 2 in. in length occupying the entire center sections of each specimen. The specimens were elec-trolytically polished using an electrolyte of 300 cc methyl alcohol, 60 cc H2SO4, 130 cc HCl, and a current density of 4 amp per sq in. Each specimen was then loaded in a bending apparatus similar to that used by Yen and Hibbard,&apos; as shown in Fig. 1. The load was applied by means of a spring and screw arrangement through four ball-bearing surfaces set 1.5 in. apart. Either the inner two bearings or the outer two (A, B, Fig. 1) were attached to a movable steel plate (D) through which the load was transmitted. Since the apparatus was mounted on a track of the X-ray apparatus, the tension or compression side of the specimen could

Jan 1, 1954

By G. Sachs, G. Espey

IT is well known that high residual stresses are created in tubing by the sinking process, in which no internal tool or mandrel is used.1-4 In this process, the wall thickness is usually slightly increased.4,5 In tube drawing with a plug, the residual stresses are gradually reduced with increasing reduction of wall, and, by the conventional approximation methods of stress measurement, no stress is detected in tubing in which the wall reduction is over 50 per cent of the total reduction in cross-sectional area.1-6, It has also been observed that higher stresses are set up by the use of wide-angle dies than by the use of acute dies.2 The actual stress distribution in drawn .tubing has been investigated by a few workers,7-9 but has not been related to the conditions of the drawing process: Some attempts have been made to correlate accelerated cracking tests for stresses in brass products with the magnitude of the circumferential stresses and the condition of the metal.1-10 MATERIALS AND PREPARATION The material used in this experiment was commercial cartridge-brass (70 per cent Cu, 30 per cent Zn) tubing in the straightened and annealed condition.‡ This tubing was of various outside diameters and had a wall thickness of 0.032 in. It was sunk to 1/2-in. outside diameter by means of an experimental 10,000-lb. hydraulic draw-bench, with a maximum drawing speed of 8 ft. per minute. All dies were of 1/2 -in. diameter. Various die materials and contours were used in order to study the effect of die design on residual stresses in the tubing; the specifications for the dies are given in Fig. I. Castor oil was used throughout as lubricant. Chattering occurred in sinking some of the tubing, particularly when dies having the more acute angles were used, but the tubes showing chatter marks did not differ markedly from those with a smooth surface regarding the magnitude of residual stress. The source of the chattering has not yet been revealed. The wall of the tubing becomes thicker during sinking, as previously observed 4,5 The increases in wall thickness were 3, 7 and 9.5 per cent on the average for 6, 20 and 33 per cent reductions in diameter, no effect of the contour being noticeable. The reductions mentioned throughout this investigation do not consider this fact, but are reductions of the outside diameter. For the evaluation of the residual stress two fundamentally different methods were used. The approximation method1,11 of splitting a certain length of tubing was employed to obtain a bird&apos;s-eye view of the effect of various drawing conditions on the circumferential stress measured by the diameter change occurring on splitting. For the determination of the complete stress

Jan 1, 1941

By R. C. Buehl, J. P. Riott, E. P. Shoub

PILOT-PLANT tests have demonstrated that it is possible to produce low-sulphur sponge iron (0.03 to 0.05 per cent sulphur) as a continuous process in an internally fired rotary kiln from iron ore or mill scale; a solid carbonaceous reducing agent; and dolomite to control the sulphur content. The proportion of dolomite employed is small, so that it does not add materially to the cost of the sponge iron. The entire operation is carried out at temperatures well below the melting point of iron, so that the product is discharged in granular form. Compared with other methods for the production of sponge iron, the rotary-kiln method has the advantage of simplicity and low cost of equipment and; in addition, requires little labor for the operation of the process. The pilot-plant equipment used for this work was sufficiently large to demonstrate that the production of low-sulphur sponge iron in rotary kilns is feasible and that the costs are not excessive for applications in which sponge iron has distinct advantages over competing materials. This report constitutes a part of a program by the Bureau of Mines under a special Congressional appropriation for studying gaseous and solid fuel reduction of iron ores. Numerous articles on various phases of these investigations have appeared in Bureau of Mines publications or technical journals. Other articles are being prepared and will be published in the near future. The present program, which is under the supervision of Dr. R. S. Dean, Assistant Director of the Bureau of Mines, was initiated in .1942. However, a considerable portion of the work was a continuation of previous research performed by the Bureau. Previous investigators1-4 utilizing the rotary kiln for the production of sponge iron did not develop a suitable method for controlling the sulphur content. A review of the literature shows that the sulphur content in the product usually exceeded 0.2 -per cent unless reducing agents with very low sulphur contents were employed. This high sulphur value made the product generally undesirable for use as melting stock in electric or open-hearth furnaces. The production of sponge iron in an internally fired rotary kiln has been described in detail in Bureau of Mines Bulletin 2702; consequently, this article will be confined mostly to new features of the process-that is, means for controlling the sulphur content of the sponge iron. Suitable raw materials include iron ores or mill scale; coal or coke for reducing them, and sized dolomite (-20+100-mesh) as the sulphur-controlling agent. A mixture of the raw materials in granular form is fed into the elevated end of the kiln, which is fired axially from the lower or discharge end. In the reducing zone, which is approximately one third the length of the kiln measured from the dis-

Jan 1, 1946

By E. P. Shoub, J. P. Riott, R. C. Buehl

Pilot-plant tests have demonstrated that it is possible to produce low-sulphur sponge iron (0.03 to 0.0; per cent sulphur) as a continuous process in an internally fired rotary kiln from iron ore or mill scale; a solid carbonaceous reducing agent; and dolomite to control the sulphur content. The proportion of dolomite employed is small, so that it does not add materially to the cost of the sponge iron. The entire operation is carried out at temperatures well below the melting point of iron, so that the product is discharged in granular form. Compared with other methods for the production of sponge iron, the rotary-kiln method has the advantage of simplicity and low cost of equipment and, in addition, requires little labor for the operation of the process. The pilot-plant equipment used for this work was sufficiently large to demonstrate that the production of low-sulphur sponge iron in rotary kilns is feasible and that the costs are not excessive for applications in which sponge iron has distinct advantages over competing materials. This report constitutes a part of a program by the Bureau of Mines under a special Congressional appropriation for studying gaseous and solid fuel reduction of iron ores. Numerous articles on various phases of these investigations have appeared in Bureau of Mines publications or technical journals. Other articles are being prepared and will be published in the near future. The present program, which is under the supervision of Dr. R. S. Dean, Assistant Director of the Bureau of Mines, was initiated in 1942. However, a considerable portion of the work was a continuation of previous research performed by the Bureau. Previous investigators1-4 utilizing the rotary kiln for the production of sponge iron did not develop a suitable method for controlling the sulphur content. A review of the literature shows that the sulphur content in the product usually exceeded 0.1 per cent unless reducing agents with very low sulphur contents were employed. This high sulphur value made the product generally undesirable for use as melting stock in electric or open-hearth furnaces. The production of sponge iron in an internally fired rotary kiln has been described in detail in Bureau of Mines Bullelin 2702; consequently, this article will be confined mostly to new features of the process—that is, means for controlling the sulphur content of the sponge iron. Suitable raw materials include iron ores or mill scale; coal or coke for reducing them, and sized dolomite ( — poi-100-mesh) as the sulphur-controlling agent. A mixture of the raw materials in granular form is fed into the elevated end of the kiln, which is fired axially from the lower or discharge end. In the reducing zone, which is approximately one third the length of the kiln measured from the dis-

Jan 1, 1948

By R. C. Buehl, J. P. Riott, E. P. Shoub

Pilot-plant tests have demonstrated that it is possible to produce low-sulphur sponge iron (0.03 to 0.0; per cent sulphur) as a continuous process in an internally fired rotary kiln from iron ore or mill scale; a solid carbonaceous reducing agent; and dolomite to control the sulphur content. The proportion of dolomite employed is small, so that it does not add materially to the cost of the sponge iron. The entire operation is carried out at temperatures well below the melting point of iron, so that the product is discharged in granular form. Compared with other methods for the production of sponge iron, the rotary-kiln method has the advantage of simplicity and low cost of equipment and, in addition, requires little labor for the operation of the process. The pilot-plant equipment used for this work was sufficiently large to demonstrate that the production of low-sulphur sponge iron in rotary kilns is feasible and that the costs are not excessive for applications in which sponge iron has distinct advantages over competing materials. This report constitutes a part of a program by the Bureau of Mines under a special Congressional appropriation for studying gaseous and solid fuel reduction of iron ores. Numerous articles on various phases of these investigations have appeared in Bureau of Mines publications or technical journals. Other articles are being prepared and will be published in the near future. The present program, which is under the supervision of Dr. R. S. Dean, Assistant Director of the Bureau of Mines, was initiated in 1942. However, a considerable portion of the work was a continuation of previous research performed by the Bureau. Previous investigators1-4 utilizing the rotary kiln for the production of sponge iron did not develop a suitable method for controlling the sulphur content. A review of the literature shows that the sulphur content in the product usually exceeded 0.1 per cent unless reducing agents with very low sulphur contents were employed. This high sulphur value made the product generally undesirable for use as melting stock in electric or open-hearth furnaces. The production of sponge iron in an internally fired rotary kiln has been described in detail in Bureau of Mines Bullelin 2702; consequently, this article will be confined mostly to new features of the process—that is, means for controlling the sulphur content of the sponge iron. Suitable raw materials include iron ores or mill scale; coal or coke for reducing them, and sized dolomite ( — poi-100-mesh) as the sulphur-controlling agent. A mixture of the raw materials in granular form is fed into the elevated end of the kiln, which is fired axially from the lower or discharge end. In the reducing zone, which is approximately one third the length of the kiln measured from the dis-

Jan 1, 1948

By Donald W. Scott, Adam Wesner

This paper describes the study of 23 nonmagnetic iron-formation samples from the Mesabi Range and shows the significance of their chemical, mineral, and physical properties in terms of their concentration. Data presented here resulted from a research project carried out from 1942 to 1952 by 10 major producers of iron ore, working in conjunction with Battelle Memorial Institute. THE inevitable depletion of Mesabi direct-shipping iron ores, hastened by recent wars, has increased the importance of developing methods to concentrate iron-formation materials. Many mining companies, independent research organizations, chemical companies, and equipment manufacturers are devoting attention to this important work. It is only natural that magnetic iron-formation materials have received major emphasis, since magnetic recovery methods apply and reserves of Mesabi magnetic taconites are large. However, the nonmagnetic* iron-formation materials are likewise believed to represent a large reserve and their study has not been neglected. tions over the years have shown that in addition to the unenriched taconite, which is dense and hard, and the direct-shipping ore, there are transition taconites grading into the friable wash, or intermediate ores. The transition taconites are less massive than the unaltered taconites and usually coarser-grained. This variation, coupled with the fact that there were no extensive drillings in the unaltered taconite, posed a most difficult problem of obtaining samples that would be representative of large tonnages of the fresh rock. The problem was increased by the tremendous extension of the nonmagnetic formation materials on the Mesabi and by the question of their accessibility to mining. A drilling program was evolved that would surmount these difficulties. Selection of samples was made by a committee representing producers in the area. The samples were procured from 1—taconite walls near bodies of stripping burden or wash ore and 2—from shafts in the process of being sunk through taconite for drainage or mining operations. Most of the samples were obtained from the cherty horizons because it became obvious early in the program that the slaty horizons represented an exceedingly difficult concentration problem, owing to the character of the gangue minerals and the fine grind required for liberation of the valuable minerals. More samples were procured from the lower cherty horizon than from the upper cherty because it was more accessible to mining. It is not known how well the samples selected represent the nonmagnetic iron-formation material. However, they were chosen by people Well acquainted with the geology of the Mesabi range and the mining problems involved. It is believed that they represent a fair cross-section of the types of materials that will be encountered. Table I gives data on the origin of the 23 samples and the horizon footages from which they were obtained. Samples 1 through 10, and 15 and 16

Jan 1, 1955

By W. H. Brandt

ThEre has been much progress in the last two decades in understanding the hardenability of steel. Roughly, the progress has been along two lines, which may be designated as empirical and fundamental. This empirical work has progressed to the point where steels may be manufactured to the desired characteristics quite closely and heat-treatment and application problems can be approached with considerable confidcnce. Particularly noteworthy among the empirical work is that of Grossmann,12. who has given formulas for calculating hardenability as a function of composition, and the establishment of the isothermal transformation curve by Davenport and Bain.13 The latter does not fall entirely in the class of an empirical study, for while it provides a means of presenting a tremendous amount of empirical data in the form of a single curve, it also outlines clearly the problems that face the fundamental investigator. The isothermal transformation curve shows clearly that the phenomena occurring in the neighborhood of the knee of the curve are those that limit the hardenability of steel. Above the knee of the curve, the product of decomposition of austenite is lamellar pearlite and fundamental progress has consisted chiefly of a detailed study of structure and formation of pearlite, largely at the hands of Mehl and his collaborators.&apos;-5 They have shown that formation of pearlite proceeds by a process of nuclea-tion and growth and they measured both the nucleation and growth rate. However, there is still a wide gap between our empirical knowledge and fundamental knowledge, owing largely to the fact that there is no satisfactory theory of nucleation rate or of growth rate. This paper is being written in an attempt to Hl in a small part of the gap. A method was prcsented recently1&apos; for calculating the velocity of edgewise growth of pearlite using an expression for the carbon concentration that satisfies the diffusion equation. The edgewise velocity of pearlite growth in a plain carbon eu-tectoid steel was calculated, using linear extrapolations of the ferrite and comentite solubilities. While the results are in fairly good agreement with the experimental facts, the calculated velocity of edgewise growth is smaller at 600°C. than G, the radial growth velocity of pearlite nodules, whereas one would expect it to be larger. In this paper the calculation will be repeated using a new extrapolation for the ferrite solubility, which gives better agreement between G and T&apos;, the calculated velocity of edgewise growth. A calculation will also be made of the effect of carbon content on velocity of edgewise growth, and the problems of calculating the effects of other alloying elements on growth rates will be discussed. It will be assumed that the mechanism of pearlite growth, established by Mehl and

Jan 1, 1947