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By S. A. LoPilato, J. T. Smith

SEVERAL techniques have been evaluated for cooling the throat area of rocket-nozzle inserts to prevent erosion or fracture of the inserts during exposure to high operating temperatures and pressures. Self-cooled composites, such as infiltrated tungsten, have shown considerable promise for these applications. These composites can be fabricated by infiltrating the refractory matrix with a lower-melting metal or non-metal which is insoluable in tungsten. During firing, the lower-melting substance undergoes a series of continuous endothermic phase changes which absorb sufficient heat to maintain the nozzle matrix material at or below a safe operating temperature. At present, the state-of-the-art has reached a point where the refractory matrix density and microstruc-ture have been established as the critical parameters to be controlled for optimum self-cooled performance.'" Nondestructive testing has been utilized to ensure that porous tungsten billets were of uniform density and free from cracks or other defects prior to fabrication into rocket nozzle inserts. Ultrasonic-velocity techniques have been employed because of their reliability and reproducibility. The purpose of this note is to present and interpret the test data developed for porous tungsten matrices. For evaluation purposes, tungsten powders with Fisher subsieve sizes of approximately 4 and 18 p have been obtained for study from four commercial sources: the General Electric Co., Wah Chang Corp., Fansteel Metallurgical Corp., and Kennametal, Inc. Porous tungsten matrices were fabricated from the powders by hot pressing in induction-heated graphite die assemblies. Typical billets ranged from 3 in. diam, 2 in, high to 5 in. diam, 5 in. high. The ultrasonic-velocity measurement technique employed was through transmission, pulse echo at a frequency of 1.0 mc using matched barium titanate crystal transducers. The testing equipment included a pulsed ultrasonic generator, an amplifier, a specimen holder, and an oscilloscope, as shown in Fig. 1. The porous tungsten thickness was measured to i 0.0005 in, with precision micrometers while the oscilloscope achieved 0.02 psec time measurement accuracy. The time for a sound wave to pass through the billet, the bulk wave velocity, was calculated to 0.5 pct using Eq. [l]: V-} where V = bulk wave velocity, in. per psec, T = specimen thickness, in., and / = time, psec. To develop inspection standards, the bulk density of a porous tungsten billet was determined by the ASTM water-immersion technique.3 The ultrasonic-velocity measurements were completed at seventeen positions for each billet and averaged to develop a characteristic standard velocity for a particular tungsten powder at several density levels. A plot of ultrasonic velocity vs density is presented in Fig. 2 for six different tungsten powders. This graph has been established as

Jan 1, 1967

By T. Sugawara, H. Sekimoto

Electrorefining of low-grade copper alloys containing various amounts of nickel, lead, antimony, bismuth, arsenic, selenium, silver, and gold in sulfamic acid media was demonstrated. The copper concentration of the alloys used as an anode was 70–90 wt%. The electrolyte contained 100 g L-1 of free sulfamic acid and 40 g L-1 of cupric ion. Galvanostatic electrolysis was conducted at 200–300 A m-2 in the electrolyte using a low-grade copper alloy anode and a stainless steel cathode at 25C. The terminal voltage was approximately 0.8 V at the start of the electrolysis, although it increased periodically to approximately 2.0 V and then shortly returned to its initial value. Nickel, lead, and arsenic from the anode dissolved in the electrolyte and were removed. Antimony, bismuth, selenium, and silver were removed as solids. Consequently, a dense copper cathode was obtained, and the concentration of copper was evaluated by inductively coupled plasma – optical emission spectrometry to be 99.82–99.99 wt%. INTRODUCTION Electrorefining of copper is conventionally carried out in a sulfuric acid electrolyte using crude copper assaying at 99.5 wt% Cu as the anode (Biswas & Davenport, 1994). Elements more noble than copper do not react anodically, and consequently, they are distributed in anode slimes. Elements less noble than copper dissolve from the anode. Nickel and arsenic are stabilized as aqueous species. Lead, antimony, and bismuth react with water, sulfuric acid, or arsenic acid to form barely soluble compounds; thus, they are distributed in anode slimes. An emerging issue in the electrolytic refining of copper is an increase of impurities in crude copper used as the anode (Moats et al., 2013). This is due to an increase in the content of impurities in copper concentrates as well as an increase in the types of secondary copper feedstocks (e.g., E-Scraps, automotive shredder residues, and smelting residues) treated in the copper pyrometallurgical process. An extreme case is a copper smelter that does not use copper concentrates as feedstocks. In such a process, copper-bearing feedstocks are smelted in a top submerged lance furnace to prepare low-grade copper alloys assaying at 80– 90 wt% Cu (Sekimoto et al., 2013). Such low-grade copper is not acceptable as an anode in the conventional copper electrolytic refining process because it is easily passivated during electrolysis. Consequently, electrolytic copper is produced by leaching the atomized low-grade copper in a sulfuric acid solution with oxygen, followed by electrowinning. However, electrowinning should be avoided whenever possible because the electrical energy consumed in electrowinning is 5–10 times higher than that consumed in electrorefining. Therefore, the development of technologies for electrorefining of low-grade copper alloys is desirabl.

Jan 1, 2019

By M. J. Saarivirta, P. P. Taubenbla

This paper presents some high-temperature properties of copper-zirconium conductor alloy compared to copper-chromium alloy. Definite superiority of the copper -zirconium alloy over the copper-chromium alloy at high temperatures is shown. RECENT investigations have shown that cold-worked copper-zirconium alloy possesses good room-temperature strength, ductility, and conductivity and maintains most of its strength at temperatures up to 450°C (840°F). As a result, this alloy is now used for many applications, such as commutator segments, resistant welding wheels and tips, rectifier bases, electrical contacts, and so forth. Early investigations1-3 have shown that the maximum solid solubility of zirconium in copper is close to 0.15 pct and the alloy containing Cu-0.10-0.15 pct Zr exhibits the optimum mechanical and physical properties. Results2 have also shown that the alloy has better elevated temperature strength and ductility than copper-chromium and copper-silver alloys. The present investigation was undertaken to learn more about the elevated temperature properties of copper-zirconium alloy and to compare these properties with those of the copper-chromium alloy. WORK PROCEDURE The commercial copper-zirconium alloy, Amzirc was selected for this investigation. Because this alloy is composed of oxygen-free high-conductivity copper and 0.10-0.15 pct Zr, it was believed that possible disturbance from impurities would thus be negligible. Copper-chromium alloys used in this investigation contained 0.40 and 0.70 pct Cr and balance copper. The Cu-0.40 pct Cr alloy was a commercial brand, whereas the Cu-0.70 pct Cr alloy was made from OFHC copper and Cu-5 pct Cr master alloy. It was made in a graphite crucible under argon gas cover and cast. In the preparation of specimens, all three alloys were treated to give the optimum properties; thus the Cu-0.15 pct Zr alloy was solution annealed at different temperatures to determine the desirable solution annealing temperature. It was then cold worked and aged at 375" to 400°C (705" to 750" F). A solution annealing temperature of 1000°C (1830°F) was selected for the copper-chromium alloys because this temperature is widely used in commercial practice. The same alloy was aged at 450°C (840°F) after solution annealing and cold working. It was found that the maximum precipitation hardening of the copper-chromium alloy occurred at this temperature. When aged after solution annealing, the aging temperature was 500°C (930"F) for both alloys. The following properties were determined: 1) Tensile properties as a function of temperature. 2) Effect of solution annealing temperature on high-temperature tensile properties. 3) Stress to rupture strength at 350° and 400°C (660"to750°F). 4) Impact strength. 5) Electrical conductivity and resistivity as function of temperature. 6) Thermal expansion.

Jan 1, 1961

By Arthur F. Taggart, T. C. Fitt, A. W. Thomas

It was noticed early by operators that high recoveries and flocculation of the sulphide minerals were closely correlated in agitation-froth flotation. Later, this readily visible flocculation was found almost invariably to be due to the binding together of the oiled sulphides by minute air bubbles, and is now known to be characteristic of pulps in which at least one of the mineral species is coated with an insoluble collecting oil. In 1930 it was reported1 that in differential flotation a correlation exists between Brownian movement of particles in the pulp and their nonfloatability, and that Brownian movement and dispersion are likewise correlated, whence it would appear to follow that dispersion of a mineral particle in a flotation pulp denotes a particle condition unfavorable to flotation. In 1934 it was postulated2 that Brownian movement, dispersion and nonfloatability all flow from the condition of the surface of the dispersed particles. It was also postulated at that time that the surfaces of dispersed particles are ionized, the ionized compound being a product of reaction between a component of the particle and a component of the pulp, of a mass solubility of the general order of a fraction of a milligram to I 5 to 20 mg. per liter. It was then further asserted, and has since been confirmed by Bankoff,3 that sulphides so conditioned with soluble collectors as to be floatable flocculate without the intervention of air bubbles as binding media. The present paper comprises an inquiry into the nature of the bound and counter ions in zinc sulphide dispersed by sodium arsenate. A part of its findings was mentioned in the 1934 paper above cited.= The tests show that: I. When powdered granular zinc sulphide is immersed in aqueous solutions of sodium arsenate, arsenate ion is abstracted from solution and anchored to the solid surface, while sulphate and lower sulphoxy ions are concurrently displaced therefrom in substantially stoichiometric quantities. 2. The ion counterbalancing the anchored arsenate ion is zinc. 3. Dispersion and Brownian movement of zinc sulphide are concurrently maxima at an arsenate-ion concentration lying between 0.01 and 0.001-molar sodium arsenate. 4. Any condition or reaction that tends. to remove the zinc arsenate coating from the zinc sulphide particles decreases dispersion and Brownian movement concurrently. Experimental Three samples of zinc sulphide were used. Lot I was C.P. zinc sulphide, washed well be~ond absence of chloride, as shown by test with silver ion. Lot 2 was prepared by precipitation from hydrochloric acid solution of arsenic-free zinc, and was washed as lot I. Lot 3 was prepared by heating a portion of lot 2 in absence of air at 1200° to r300°C. for one hour."

Jan 1, 1943

By A. Kenneth Schellinger, O. Cutler Shepard

Overgrinding, Or the breaking of ore particles into sizes smaller than required for liberation, is a first-magnitude problem in grinding for concentration processes. The conventional ball mill-classifier circuit used in commercial grinding produces mineral particles that are all ground to pass an upper size limit, while the proportion of very fine sizes, or slimes, is uncon rolled. Calculations based on measurements made in commercial plants which were grinding ore for flotation concentration indicate that about go pct of the energy used in grinding is consumed in useless overgrinding.1 In addition to accounting for the consumption of a large amount of energy, over-grinding is detrimental to the flotation process. Overground slime coats larger partitles and inhibits selective action by the flotation reagents on particles which by themselves could be separated very well. A further difficulty is that extremely finely ground minerals do not respond satisfactorily to flotation-collecting reagents, and both selection and recovery decrease markedly with particles finer than 5 to 10 microns.2 conventional classifier-ball mill circuits inherently tend to overgrind the valuable minerals more than the gangue minerals because most valuable minerals are of higher density and are softer and more friable than the gangue minerals. The rem- edy to this situation would seem to be a separating mechanism which would separate free valuable mineral particles regardless of size and density considerations and a mechanism which would operate in the grinding mill so that valuable mineral grains would be removed from the grinding machine as soon as they are freed. A consideration of these ideas led the authors to the conception of a simultaneous grinding and flotation mechanism. So far as the authors are aware, such a machine has never been tried, even on a laboratory scale. Development oF Apparatus and Experimental Method Apparatus development went through several stages before a really workable machine was devised. Work was started with a large cast-iron mortar 7½ in. in diameter and 8 in- deep. A cast-iron disk Was mounted on the end of a vertical shaft SO that it fitted loosely into the bottom of the mortar. On top of the disk four radial cleats ¼ in. high were bolted to agitate the ball charge as the disk, or impeller, rotated. The vertical shaft was driven at 100 rpm by a mechanism improvised from an old coffee-mill type of grinder. A hole was drilled into the bottom of the mortar through which compressed air was introduced for frothing. In operation, more charges of Ore which had been crushed to pass an 8-mesh screen were placed in the bowl along with water and suitable reagents. The impeller was then started and the compressed air turned On so that the ore was ground while being exposed to air bubbles. Mineralized froth was

Jan 1, 1949

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 R. W. Gurry

Thermodynamic calculations indicate that boron is a better deoxidizer than silicon but probably is not quite as effective as aluminum. Boron should, therefore, be readily oxidized out of the open-hearth bath, which implies that it is not recovered in significant concentration from any scrap charged; also that, if added in the ladle, it tends to oxidize out during teeming unless protected by the presence of a better deoxidizing agent. These conclusions do not apply to liquid cast iron, in which the oxygen content is much lower than in liquid steel. The possibility of conserving certain strategic alloying elements widely used in steelmaking, notably manganese, chromium and molybdenum, by the use of boron, which in small concentration appears to be extraordinarily effective in promoting hardenability, raises a question as to what extent any boron present in the scrap charged into the open hearth is oxidized out during melting and subsequent refining. This question is significant because the optimum concentration of boron in the finished steel, which is fairly critical, would be very difficult to attain if the boron in the charge were largely retained in the metal. Moreover, unless boron is substantially eliminated in the open-hearth process its use on a large scale would eventually make it impossible to produce boron-free steel except by restricting the charge to virgin material. There is at present, unfortunately, virtually no satisfactory direct evidence to provide an answer to this question; but there are indications from practice that boron is oxidized out in the open hearth in the production of medium or low-carbon steel, although it may be retained if the carbon content remains fairly high; that is, if the oxygen is kept fairly low. In the absence of direct data, it is necessary to rely upon the indications of indirect evidence, the best approach now available being a thermodynamic calculation of the deoxidizing power of boron in liquid steel by the general method used so successfully by Chipman1 in his study of the reactions in the open-hearth process. The thermal data for boron, upon which such a calculation must be based, are reasonably satisfactory; the only quantities not directly determined are the entropy of melting of boron, the activity coefficient of boron in liquid steel, and the activity coefficient of B2O3in an open-hearth slag; but these values may be estimated with ample accuracy. The results of this calculation, which is presented later, are given in Table I, together with corresponding data for other deoxidizing elements. This table shows the calculated residual content of oxygen in equilibrium with different residual concentrations of the deoxidizer in the metal;

Jan 1, 1944

By Arthur Phillips, E. N. Skinner

DURING the course of an experimental program concerned with the hydrogen embrittlement of copper containing oxygen in concentrations within the solubility limits it became necessary to make a critical appraisal of the literature dealing with the solubility relationships of oxygen and solid copper. The latest determinations, and perhaps the only data obtained on "spectroscopically" pure copper, are those of Rhines and Mathewson1 who reported that the solid solubility increases from 0.007 per cent oxygen at 600°C. to about 0.015 per cent at 1050°C. Hanson, Marryat and Ford2 had previously placed the solubility limit at less than 0.009 per cent oxygen at 1000°C. and Allen and Street3 had suggested, on the basis of an unpublished research by T. Hewitt, not over 0.005 per cent oxygen at 500°C. In the introductory stages of the investigation indirect evidence accumulated to suggest that the solubility limits were significantly lower than those previously reported. For this reason it was deemed essential to clarify this point before proceeding with a program intended to establish a quantitative relationship between hydrogen embrittlement and oxygen concentrations positively known to be within the solid solution range. Furthermore, the increasing production and the highly selective applications of the commercial "oxygen-free" coppers inevitably direct attention to the practical necessity for exact knowledge regarding the effects of minute concentrations of the elements ordinarily found in the high-purity coppers now available. Of these elements oxygen occupies a unique position, not only by virtue of its specific effects on copper but more particularly in view of its important reactive tendencies toward the other impurities with which it is associated. Rhines and Mathewson considered several methods for measuring the solubility of oxygen and finally adopted the simple plan of saturating the specimens, obtained from redeposited cathode copper, by heating at the desired temperature in air. After the removal of surface oxide by mechanical and chemical means, the oxygen content was determined in the underlying metal. In general, this process of saturation, with a few modifications, was followed by the present writers. Although the procedure is direct and has obvious advantages, its effectiveness, unfortunately, is reduced by an excessive scaling of the copper, particularly at the higher temperatures. After saturation treatments above 900°C. the writers found that the specimens, necessarily rather thin in the original condition, consisted largely of scale and yielded too little metal for a satisfactory analytical sample. Furthermore, great difficulty was experienced in effecting complete removal of the scale. Needless to say, it is necessary to exercise extraordinary care in the preparation of the surface of the analytical sample in order to eliminate submicroscopic fragments of cuprous oxide which ordinarily are lodged in the crevasses and

Jan 1, 1941

By D. H. Davis

HAND picking was the earliest form of coal preparation, first practiced to improve the outward appearance of the coal being loaded and to remove any pieces that might appear objectionable to the buyer. It is still extensively practiced on the plus 1-in. sizes and particularly on the plus 4-in. and larger sizes. The percentage of ash reduction that can be obtained by hand picking depends entirely upon the amount of free impurity occurring in the coal. In a particularly dirty seam, where a considerable quantity of rock occurs with the coal as mined, hand picking produces sizable reductions in ash. If the coal is prepared clean at the working face hand picking is a final check; in some cases, on the smaller sizes, it is practiced only to show that an effort is being made to prepare a clean product. Because of the variety and different methods of handling coal, the early plants were not always adapted for e5cient hand picking. At most mines an attempt is made to - remove the visual impurities at one or more of the following places in the surface plant: (1) in the tipple during screening; (2) on picking band, belt, or shaking conveyors; (3) during the loading of the coal as it goes into the railroad car. Whatever the physical arrangements of the plant may be, the usual practice is to run the coal over the screen, picking table or boom at such a rate that each piece of impurity is exposed to the view of the picker. Coal should be uniformly sized and spread out on the table so as not to exceed one particle in depth. The flow should be even in tonnage and the speed of travel should be from 30 to 60 ft. per minute and not exceed 80 ft. per minute. The dirtier the coal, the slower the speed required for a fixed number of pickers. Actually, picking capacity depends on the number of particles per minute that can be picked out and rejected. The percentage of impurity removed is determined by this and the average weight of piece rejected. Handling a mixture of sizes from nut to the largest lump over the same picking table or conveyor is not conducive to efficient separation by hand picking because the lump hides the smaller impurities. Picking tables should not exceed 5 ft. in width, so that a picker on either side will be able to remove impurities from the center of the table. Guards usually are placed at the side of the screen, boom or table, to allow the pickers to stand close enough to remove the impurities without too long a reach.

Jan 1, 1950

By David Williamson

In the post-World War II years, metals analysts and statisticians have tracked the demand for metals in line with changes in Organization of Economic Cooperation and Development (OECD) industrial production (IP) growth rates, Williamson said. This was because the correlation between IP growth and metal demand was reasonable. "However, the correlation was exaggerated," he said. During times of expanding IP growth, metal demand increased at a faster proportional rate than IP growth. This was probably due to industry restocking in anticipation of growing demand at the start of each economic cycle. Conversely, metal demand would slacken at a greater rate than would be suggested by negative IP growth as industry destocked to reduce costs. Nonetheless, he said, a reasonable correlation could be painted between IP growth over the longer term and metal demand. "Yet, although this correlation has stood well since the last world war, it has been based on the fact that the bulk of the world's IP growth stemmed from the industrialized countries of the western world, lately including Japan," Williamson said. The so-called "developing world" had been of little consequence. The Soviet Union and other communist bloc countries were deemed to be closed, self-contained entities that, for each metal, would either be net exporters or importers, he said. These quantities would be used by analysts to calculate the overall supply/demand equation for the western world only and to deduce whether a commodity was in shortage or surplus. It is now believed that this method of projecting total world supply and demand for raw materials is outdated, Williamson said. It does not take into account the worldwide political changes that have occurred since the collapse of communism and the revitalization of China. "It is no longer appropriate to assess global balances for raw materials on the basis of three economic blocs," he said, including the western world, developing world and communist bloc. Rather, global balances need to be projected on a global basis. In this regard, since the collapse of communism and the removal of many other totalitarian regimes, capitalist forces are attempting to satisfy the expectations of the great mass of the world's poorer people, Williamson said. Such people, usually located in Asia, Africa and Latin America, have aspired to the same standard of living as people living in western Europe, North America and Japan. "Now, with free market forces at work, their aspirations will start to be met." Consequently, the demand for metals and other raw materials will now rise at a much faster rate. The production of raw materials has to keep pace with rising IP in the developed world. It will also have to keep pace with the demands associated with the rising expectations of increasing populations in the developing world, as well as from within the old communist bloc countries.

Jan 1, 1997

By R. S. French, W. R. Hibbard

FOR tensile deformation, if the stress value is defined by the ratio of the load to the actual area, and the strain value by the natural logarithm of the ratio of the immediate length to the original gauge length of the sample, the resulting data when plotted with logarithmic coordinates result in a linear relation extending from the initial plastic yielding through the maximum load and in some cases, such as for copper and brass, to fracture'.'. The normal form of the curve, therefore, is a parabola, satisfying the following equation: S = Kem [1] where S and e represent true stress and strain, K is a constant, and m is a coefficient evaluating the slope of the curve and designated therefore as the strain hardening coefficient. Hollomon² has reported that this coefficient is a function of the yield strength as effected by grain size in alpha brass, and the carbon content and heat treatment of steel. In the case of steel, the value of the proportionality constant is a function of the carbon content. It was further shown in that paper2 and also by Low and Prater -hat, if the power relation between stress and strain remains valid, the strain hardening coefficient m, is equal to the strain at the maximum or necking load. Thus, an added significant feature is that this coefficient may serve as an index of the capacity for deformation prior to localized necking and subsequent fracture, e.g. drawability. The constant K may be defined as the strength of the material at unit strain. Together the K constant and the m coefficient describe an alloy's physical properties as far as strength and ductility are concerned for a particular yield strength and, in the case of steel, a particular carbon content. The work of Lacy and Gensamer4 gives data on alloy ferrites from which very close estimates of the strength of ternary and higher iron alloys have been made. It was the purpose of this investigation to develop similar data for commercial copper base alpha solid solution alloys. Preparation of Specimens Commercial binary alloys of copper with aluminum, beryllium, cadmium, nickel, silicon, tin and zinc were investigated. In addition, laboratory castings were made of a 1 pct silicon alloy and a 2 pct aluminum alloy to complete these particular series. Two types of copper were used, oxygen free OFHC and tough pitch FC. The analyses of these alloys are shown in table I. Standard ASTM sheet tensile specimens were used to determine the usual physical properties of each alloy and to obtain stress data at strains from 0.0005 to 0.01. Small (1/4 in. x 3 in. x tk.) samples

Jan 1, 1951

By G. G. Gunther, C. L. Boeke

DISCUSSION I. M. Symonds, (Cia. Minera de Penoles, Monterrey, N. L., Mexico)—Messrs. Boeke and Gunther have written most interesting papers describing their zinc de-leading operations. Mr. Gunther was kind enough to show me his operation last fall. Aside from the excellent results, I was most impressed by the ease of control. I had expected a difficult operation requiring careful control and supervision, but found one requiring little attention. The only disappointing fact of the separation is the poor copper recovery in the concentrate of the delead-ing float. The final zinc concentrate still has about 3 pct chalcopyrite, assuming the copper is chalcopyrite, and adversely affects the zinc grade. If the copper mineral is free, the poor recovery is probably due to the high cyanide concentration. Fortunately, about 75 pct of the copper is recovered before the pulp goes to the zinc circuit. Mr. Gunther's remarks about proper cyanide to zinc sulphate ratios cannot be overemphasized. Although our experience has been confined mostly to the conventional lead-zinc separation, we have found these ratios most important, and more important in the mill than in the laboratory. The laboratory appears to give optimum results over a wider range of reagent control than the mill, especially in regard to the effect of cyanide. In handling relatively high grade ores in the mill we usually find that ratios of cyanide to zinc sulphate of 1 to 10 or 20 or higher work best. Where there is a substantial amount of chalcopyrite present it has been noted that cyanide in the range of 0.02 to 0.10 lb per ton of ore is often beneficial in zinc rejection and possibly in copper recovery in the lead concentrate, while larger amounts, say 0.20 to 0.25 lb per ton, will float more zinc and cause a serious rejection of chalcopyrite. On an ore with 15 to 20 pct zinc content we have had to use 0.8 to 1.2 pounds of zinc sulphate per ton of ore with the above cyanide to obtain optimum results. Some time ago we made a laboratory study following Parral and Santa Barbara practice on an ore which had never a very satisfactory rejection of zinc in the lead concentrate. The results were not encouraging and a mill test was not run. Since then the study has been reopened using in the laboratory fresh zinc concentrate from the mill instead of one made in the laboratory. A reasonably clean lead float from the zinc concentrate was obtained. It has been checked several times and each time the zinc concentrate made in the mill responds and the one made in the labora-

Jan 1, 1953

By Edward C. J. Urban

THE nature of certain pernicious dusts commonly encountered in the removal of iron ore from the underground mines of the Lake Superior district is recognized, and appropriate measures for the protection of individuals subjected to these dusts have been applied in most mines for more than to years. Undoubtedly the principal concern has been demanded by dusts that contain as a constituent silicon dioxide in its free or uncombined state, the mineral we all know as quartz or chert. Investigators1 agree that among the factors that determine injurious exposure are: the composition of the dust, the concentration suspended at the breathing level and the duration of the exposure. We have on hand data that describe the experience of several underground iron-ore mines in the Upper Peninsula of Michigan and Wisconsin in regard to these factors, which in a large measure also reflect the general experience of the other mines of the area. We believe that a discussion of these data will indicate the effectiveness of the dust-control methods that have been considered good practice and that have been employed over a long period of time. A silicosis-prevention program was ini¬tiated by the iron mines of the Upper Peninsula in 1934 and records of mine dustiness are available from that date to the present time. For dust conditions previous to 1934 the results of observations conducted in other industries can be utilized together with local information pertaining to the development and initial application of rock-drilling equipment. DEVELOPMENT OF ROCK DRILLS The earliest percussion rock drill that later had any industrial use was patented in this country in 1851.2 This machine was later modified and improved in many respects but its piston-type design remained standard for drilling equipment in most mines for the next 63 years.3 The piston drill was operated either by air or by steam. It employed a solid drill steel, fastened directly to the reciprocating piston. The drilling was done dry except in holes directed at an angle below the horizontal. For these water was manually injected into the hole and the pumping action of the rod expelled the sludge. The dry cuttings from drillings inclined above the horizontal fell from the hole by their own weight. A development of major importance from the standpoint of dust generation and control occurred with the invention by J. G. Leyner of a hammer-type rock drill that embodied the essential feature of a drill steel with a hollow core. His patent was secured in 1897. The machine was a radical departure in drilling method. The steel was held loosely in a chuck while

Jan 1, 1947

By Maurice Rey, H. Beavers

THROUGH circumstances connected with the war, the senior author lost his records, therefore it has been impossible to include numerical data in this paper. The arsenical quicksilver ore investigated had been roasted, with poor metallurgical results. The vapor-tension temperature curves of mercury and arsenious trioxide are similar, and the arsenical compound hampers the coalescence of mercury drops in the condenser and is the cause of large stack losses. The ore was a gray limestone carrying small amounts of carbon and mineralized with cinnabar and two arsenic minerals, realgar (As2S2) and arsenopyrite (FeAsS). The proportion of realgar often exceeded that of cinnabar; the proportion of arsenopyrite was smaller. The grade of the ore varied between 0.8 and 4 per cent Hg. The problem was to make a high-grade mercury concentrate, as free from arsenic as possible. The water used throughout the testing was hard, having a pH of about 8. FLOTATION OF REALGAR Realgar showed a very strong tendency to float. It formed a film at the surface of the pulp and floated in abundance with only pine oil. Therefore an attempt was made to float the realgar before the cinnabar, but this was not very successful; the percentage of mercury in the realgar concentrate could never be brought below 2 or 3 per cent. Cinnabar slimed easily and floated with pine oil, so that some of it always accompanied the realgar. There was also the question of the cinnabar-arsenopyrite separation. On further investigation it was found that realgar can be floated not only with pine oil but even better with a combination of hydrocarbons and pine oil. Washing the ore with ether revealed that it contained hydrocarbons, which helped to float the realgar. In this connection it belongs to the same class as carbon, sulphur and talc. It can be floated also with a xanthate, when activated by a heavy-metal salt. The conditions favoring the two types of flotation are opposed, as activation by a copper salt facilitates xanthate flotation (in fact, is even necessary to it), whereas it hinders pine oil and petroleum flotation. One type of flotation can be termed nonpolar and the other polar, a distinction that has been stressed in a previous paper.1 DEPRESSION OF REALGAR AND FLOTATION OF CINNABAR 'Many chemicals were tested in an attempt to float the cinnabar and depress the realgar. Finally starch rendered soluble with caustic soda was tried, also several grades of commercial soluble starch, and dextrin was found to be the best and the most convenient reagent. Several of these starch preparations have a strong flocculating effect on the pulp, which is a distinct disadvantage. Dextrin, with its low flocculating power, is excellent. It is noteworthy that starch preparations are also depressants for carbon and talc, so that the fact apparently has a general

Jan 1, 1941

By Frank H. Kingdon

and eventually we would have trouble. That is, these little localized overstresses may not be important if the structure is loaded a few times, but may be important if the structure is loaded many thousands of times. The existence of localized stress has been demonstrated by direct experimentation with delicate extensometers by the elaborate investigations of Coker. Localized stresses four, five, and six times those that we would get by the ordinary laws of mechanics exist in most of our structures. These stresses, ordinarily, under steady load do no damage but if they are repeated many times may start progressive failure. So far, we have seen no failure that could not be explained by localized high stress at a nick, a crack, a point where one heat treatment stopped and another began, or where there was local faulting, or any one of a dozen causes. In the case of drill rods, the actual distribution of stress must be different from that which any mathematical analysis could predict. It would seem that the spread of damage from the localized overstresses, so far as we can see now, explains the phenomenon known as fatigue. The Ideal Drill Steel By Frank H. Kingdon The essential qualities of a drill steel are: (I), It must be easily forged; (2), the forged bit end must permit easy heat treatment to obtain hardness to resist chipping; (3)) the bar or body must be stiff to resist bending or twisting and yet tough to resist shock and vibration, with resulting breakage, (4), the forged shank end must be easily heat treated to obtain hardness with great toughness. Minor requirements are that the bit and the shank are in alignment with the body; that the shank is of the proper shape and length and the shank collar or lugs of the proper diameter and length; that the hole is free from obstruction; that the striking end of the shank is flat and square; that the bit is of the proper shape, with the cutting and reaming edges formed full and to the required size; that the reaming edges are concentric with the axis of the steel, and that there are no sharp corners at the shoulder where the bit blends into the body. The essential qualities of a bit, are: That its shape be such that maximum cutting speed can be maintained for as great a distance as possible before wear of the gage and cutting edge reduces the speed of penetration to a point where a change of steel is made necessary; that the size or diameter of the drill hole corresponding to the gage of the bit can be maintained with the least possible reduction as the depth of the hole increases, and that the bit can be correctly and readily formed and heat treated. The following features of bit design require attention: Shape,

Jan 1, 1922

By A. U. Seybolt

DURING the course of experiments involving oxygen equilibrations with a high-purity Pd-5 at. pct Rh alloy, the appearance of a subscale was noted. Most of the heat treatments in a pure oxygen atmosphere were performed at 900°C. The presence of the subscale was noted when the oxygen partial pressure was in excess of approximately 95 mm, up to 1 atm, the highest pressure used. According to available thermodynamic information' on the free energy of formation of Rh2O3 (the only reliably reported solid oxide of rhodium), this oxide should not be stable under these conditions, nor, of course, any oxide of palladium. Extraction of the subscale from the matrix was affected by treatment with aqua regia, and the resulting X-ray diffraction powder pattern is shown in Table I. This pattern is not that of Rh2O3, nor that of any known rhodium or palladium oxide. It can be indexed on a hexagonal lattice with a. = 5.22Å and co = 6.0Å. In addition to this unknown material, there was a weak Rh2O3 pattern. An X-ray fluorescence analysis of the powder gave 3214 palladium counts per sec and 5381 rhodium counts per sec, thus indicating the likelihood of a double oxide of palladium and rhodium. In view of the fact that the above result was obtained on a single sample, it is perhaps best not to insist upon the above ratio of about 0.6, but possibly to tentatively assume 0.5 or 1/1. The small amount of Rh2O3 in the X-ray pattern might have occurred on cooling, and it seems possible that the double oxide, below some temperature level, is not as stable as Rh2O3. To examine this point, some of the extracted double oxide was heated on a sheet of pure palladium in air at 900°C, and allowed to furnace-cool over a period of several hours. The X-ray diffraction pattern of the powder showed approximately the same intensity of Rh2O3 and the "hexagonal" oxide as before, but, in addition, a rather faint pattern of another crystalline species made its appearance. This finding again suggests partial transformation of the major (double) oxide to other oxides on cooling from 900°C. Application of the phase rule shows that only one equilibrium oxide (subscale) could have been present at temperature. Vacuum-fusion, oxygen analysis of the extracted subscale showed 7.7 ± 0.7 pct O. Combining this information with that of the X-ray fluorescence analysis suggests the formula PdRhO. This compound should have an oxygen content of 7.07 pct, which is within the range of error of the oxygen analysis. However, until a more careful stoichiometric study is made, it is suggested that the formula be written

Jan 1, 1965

By P. A. Beck

ACCORDING to the traditional definition,1 re-x crystallization is a process taking place upon annealing of cold worked metals, characterized by the appearance of new strain-free grains, growing at the expense of the cold worked grains until ultimately the new crystals completely replace the cold worked structure. In metals like brass this process was observed long ago to take place simultaneously with the return of the mechanical properties to those characteristic of well annealed material. The above definition implicitly assumes that the matrix in which the new grains are growing is essentially still in the cold worked condition at the time it is absorbed by the new grains, i.e., that the energy stored in the metal during cold working is not greatly decreased by recovery, prior to recrystallization. Accordingly, the bulk of this energy is assumed to serve as the driving force for the process of recrystallization. Recently, a great deal of evidence accumulated, indicating that the traditional concept may be considerably in error. Suzuki' found that by far the largest portion of the stored energy of cold work in compressed polycrystalline copper is released in the early stages of annealing, well in advance of recrystallization. Thus, the new recrystallized grains are undoubtedly growing in a matrix which is, at least energetically, no longer in the cold worked condition, but has nearly returned to the annealed state. It is true that in Suzuki's experiments the major hardness changes due to annealing were observed to take place concurrently with the formation and growth of the new grains, but this process apparently involved the release of only a minor residual fraction of the stored energy of cold work. Confirming in this respect the views of Wood," several recent investigations by various techniques lead to the conclusion that the grains of many, if not all, metals in the cold worked condition consist of small, slightly disoriented "crystallites" or "sub-grains".* (After appreciable deformation at room temperature, the subgrain size in copper appears to be about 7.104 cm and in high purity aluminum about 2.10-4 cm.) Recent investigations which demonstrated this phenomenon most convincingly are those of Heidenreich8 with hammered high purity aluminum, of P. B. Hirsch and collaborators7,8 with cold worked polycrystalline aluminum, copper, nickel, iron, gold, etc., and of Warren and Averbach9 with a rolled copper plus 2 pct Si single crystal. In addition to the existence of slightly disoriented sub-grains, the cumulative evidence from these investigations also shows that in the cold worked condition the subgrains themselves are severely strained. With regard to the early stages of annealing, Heiden-reich's results indicate that within the subarains of hammered high purity aluminum the strain gradually disappears, without discernible migration of the subboundaries. This recovery process presumably accounts for the major portion of the energy release

Jan 1, 1953

By M. K. Yen

The distortion of crystal structures referred to as "biege gleitung",1 "local curvature"2-4 and "deformation bands"5,6 is believed to be an unavoidable characteristic of deformed single crystals. The presence of asterism on Laue photograms of Al single crystals either deformed by pure shear7 or extended in a condition relatively free from the grip effect2 indicates strong evidence substantiating this point of view. The type of localized bending described in connection with the bending test of Al single crystals by Yen and Hibbard8 could conceivably exist either microscopically or submicroscopically in all types of deformation. The purpose of this note is to extend this idea to the postulate of a hypothetical "microbending site," as applied to the mechanism of slip in a general sense. The mosiac structure, as postulated by Darwin,9 is that crystals were actually composed of small blocks which have a linear dimension 10—4to 10-6 cm on the edge and deviate from each other by an angle of about one-tenth of a degree. The following discussion is based on the assumption that when slip is propagated from one mosaic block to the adjacent one, bending of the lattice at the region of their boundary is produced in order to maintain the continuity of the gliding process along a certain slip plane. Consider a two-dimensional diagram of two ideal mosiac blocks on the cross-section perpendicular to the slip plane and containing the active direction of slip. When an external force is applied, the localized bending or "microbending site," would be generated at the region as illustrated in Fig 1A at point B. A schematic diagram of a "microbending site," generated at a mosaic boundary, is illustrated in Fig IB. It may be seen that once the curvature of the bending site reaches sufficient magnitude, the interaction force encountered by pairs of atoms symmetrical with respect to the center of the bending is equal but opposite in sign. Consequently, the overall energy required to displace the atoms near the center is approximately zero.10"12 Based upon the finding that the misalignment between two mosaics is 0.1 degree,9 the assumption is made that "microbending sites" may possibly exist in the region where one plane will have one more row of atoms than an adjacent plane. It is probable that the origin of the bending site will be where the distance between two mosaic boundaries is "2d" as shown in Fig 1A at point B, where " d" is the interatomic distance. Thus, the distance between two "microbending sites" is in the range of about 600 interplanar spacings. In the case of aluminum, the spacing between two octahedral planes is about 2.33 A. The distance between two bending sites will be approximately 2.5 X 10"5 cm. It is interesting to note that according to Woods13 these values approach closely the size of the domains of a cold-worked metal whose minimum

Jan 1, 1950

By A. E. Perkins

fractured ingot will show a complete diagonal structure. If the temperature is lowered, the equi-axed zone is larger and the pine tree growth is less. A small ingot, though, is prone to this columnar structure, even though the casting temperature is kept down. The speed of cooling after solidification is slower in the hollow ingot than in the solid, therefore the granulation, which takes place after the crystals have formed from the liquid metal, is apt to produce larger grains than in the solid ingot. This is not a detriment. The original crystal size is of much more importance. The size of the grains due to this granulation constantly varies in response to mechanical work and subsequent heat treatment. Surface weakness, as shown by seaming and checking on the exterior when working or by cracking in hardening, is in some measure the result of the surface stresses of the ingot when cast. The temperature of the ingot mold or speed of freezing of the exterior of the ingot is of great importance and much seamy and cracked steel is caused by a too quickly cooled skin. Ingots that are turned or milled all over will not seam very much in working, in fact they are practically free from seams, whereas an ingot rolled without a preliminary machining operation is full of seams and cracks, some long and some short. It has been said that these short seams and cracks in the surface of the bar are the result of elongated depressions in the surface of the ingot and that if an ingot is cast very smooth it should give fewer surface defects. This is partly true, but an ingot cast very hot and having a very smooth skin will crack and seam very badly in the mill, as will the surface of an ingot poured very cold, which is very crinkly, because of the low temperature of the metal. The cause of the cracks and seams in one is not the cause of the cracks and seams in the other. The tube method of hollow drill steel manufacture is superior to the older method because of: (1) Greater freedom from external and internal straining; (2) the inherent small crystal size; (3) absence of harmful segregation, resulting in weakness of wall of hole. (4) less liability for steel to crack in inside of hole during forging or hardening; (5) toughening effect, arising from mild steel wall of hole limiting the intense hardening on quenching. Breakage and Heat Treatment or Rock-drill Steel By A. E. Perkins, Pittsburgh, Pa. Actual figures showing the superiority of one section or another are difficult to obtain. In a given size of steel, however, the cruciform sec -tion shows the most breakage; it is inherently a weak section. It contains less steel per foot than any other standard section of the same size and when rolled the wings tend to cool off setting up cooling stresses.

Jan 1, 1922

By P. D. S. St. Pierre

IN recent years the use of high sulphur fuels and charges in steelmaking has necessitated rapid methods of slag control in order to insure the production of high quality steel. Several systems of control have been discussed in the literature, and of these the method described by Philbrook, Jolly, and Henry' has received the most attention recently. In their process the hydrogen ion concentration of an aqueous slag suspension is determined under controlled, but arbitrary conditions. The basicity of the slag may then be derived from a previously established pH-V ratio curve. Thus, by measuring the pH, some indication of the V ratio is gained, from which significant properties of the slag may be deduced. Philbrook et al. made no claims as to the theoretical justification of their treatment, but simply offered it as a rapid means of slag control. Encouraging results caused others to study the method and publish their findings.2-5 Although there was fair agreement on the efficacy of the method, no common parameter was evolved against which pH readings could be plotted. The results of the various determinations were related to such V ratios as CaO/SiO2 and CaO/ SiO2 + P2O5 by some,'-3 while others" plotted against "excess lime." Rivet,' however, found a simple proportionality between the hydrogen ion concentration and the sulphur content of the molten steel beneath the slag. Since slag control in North American practice is largely directed toward the efficient removal of sulphur from steel, Rivet's approach to the problem had the advantage of connecting directly a property of the slag with the impurity being removed. In view of the diversity of parameters used it seemed probable that some fundamental point was being overlooked in seeking a relationship. Examination of the literature revealed that much emphasis had been laid on experimental detail but scant attention paid to the fundamental aspects of the hydrolysis of the aqueous slag suspension. Consequently the present investigation was undertaken with the object of rectifying this deficiency and deriving, if possible, a parameter of general application. Petrographic studies on several matured Canadian open-hearth steel slags have shown that their major constituents are dicalcium and tricalcium silicates set in a complex matrix of iron oxides and ferrites. These phases are shown in Figs. 1 and 2. Free lime is frequently encountered in slags, but is not an equilibrium phase since on continued heating it dissolves forming silicates or ferrites. Thus, it appeared that a study of the hydrolysis of the major phases in steel slags would be profitable. The experimental technique used in the present work was a modification of that described by the other workers.'- Electrical conductivity measurements were also made on the suspensions, since Smith et al." had claimed advantages for this method in their paper on slag control. Experimental Procedure The compounds studied in this investigation were prepared in the laboratory by heating purified oxides

Jan 1, 1954