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By E. R. Buchanan, R. E. Reed-Hill

Cylindrical magnesium single crystals were deformed in torsion. The axis of torsion was the basal pole. Five crystals were deformed to different strains between 0.08 and 1.03 and then sectioned in order to determine their dislocation densities using the transmission electron microscope. The work hardening rate of all specimens was at least an order of magnitude higher than that normally associated with easy glide in a tension test and comparable to that obserzed in stage B of a tension test. The correspondence between stage B in a tension test and the initial stages of a torsion test is believed to result from the fact that in both cases deformation occurs in a structure containing roughly equivalent densities of the dislocations associated with the three basal slip vectors. THERE are three basic ways of changing the shape of a crystal using slip dislocations. The crystal can be sheared along a slip direction in a slip plane; it can be bent about an axis in the slip plane normal to the slip direction; and, if the slip plane contains multiple slip directions, it can be twisted or rotated about the slip plane normal. Of the three processes, generally known as translational, flexural, and rotational slip, only rotational slip has not been extensively investigated. The object of the present research was to obtain basic information about this phase of crystalline plasticity. This paper is concerned with some aspects of the stress-strain curves obtained by twisting magnesium crystals about their basal poles. Magnesium was selected for four reasons. One, its recovery rate at room temperature is relatively low. Two, single crystals may be easily grown. Three, at room temperature the critical resolved shear stress for basal slip is about two orders of magnitude lower than that for nonbasal slip, so that rotational slip might be expected to occur primarily by basal slip. Finally, relationships between flow stress, shear strain, and dislocation density have been determined for tensile deformation of magnesium crystals,' thereby providing a basis of comparison for the torsional results. EXPERIMENTAL PROCEDURES Crystals were grown from 0.5 in. diam extruded sublimed magnesium rod furnished by the Dow Chemical Co. who reported that it contained in weight percent less than 0.0003 Al, 0.01 Ca, 0.01 Cu, 0.0012 Fe, 0.001 Mn, 0.0005 Ni, 0.003 Pb, 0.001 Si, 0.01 Sn, and 0.02 Zn, with the balance magnesium. All crystals were grown in a dried argon atmosphere using Jill-son's2 modified Bridgman technique. Two different methods were used to obtain crystals with the basal plane normal to the specimen axis. These included direct growth from seeds of the proper orientation, and spark machining (trepanning) of 0.5 in. diam cylinders from larger (1.5 in. diam) randomly grown crystals. The crystals thus produced were polished in a 20 pct HC1 water solution and then etched in a 1 pct HC1 water solution to develop etch pits lying in rows parallel to the basal plane trace. The crystals were then revolved about their axes on the stage of a low-powered microscope. Torsion specimens were prepared only from those crystals whose axes were normal to the rows of etch pits. It is estimated that this simple orientation method was accurate to about 51 deg. 1.625 in. long cylinders were next cut from these selected crystals and then acid machined in a 20 pct HC1 water solution into torsion specimens with a 1 in. long by 0.30 in. diam reduced gage section. Following this, they were annealed in air at 400°C and furnace cooled to room temperature. The specimens were given a final polish in a 10 pct HC1 water solution, rinsed in ethanol, dried, and cemented into brass grips with an epoxy resin. In addition to the torsion specimens, several tensile specimens with their basal planes at 45 deg to the stress axis were also prepared with similar dimensions except that the gage sections were increased to 4 in. in length. The torsion specimens were tested in the apparatus

Jan 1, 1970

By B. B. Rath, Hsun Hu

In wedge-shaped bicrystals of zone-refined aluminum it is observed that (111) pure tilt boundaries migrate under the driving force of their own inter-facial free energy. The boundary velocity is a power function of the driving force. The driving force exponent decreases with decreasing angle of misorien-tation. For example, at 64O°C, the exponent decreased from 4.0 for a 40 deg to 3.2 for a 16 deg tilt boundary. An evaluation of the driving force acting on the boundaries during their motion indicates that for low driv-forces, up to about 2 x l03 ergs per cu cm, the velocity is relatively independent of misorientation, whereas at higher driving forces a 40 deg tilt boundary exhibits the highest velocity. The measured activation energy for boundary migration approaches that for bulk self-diffusion at low driving forces, decreasing from 33 to 27 kcal per mole as the driving force is increased from 1 x l0 to 5 x l03 ergs per cu cm. These results are compared with current theories of grain-boundary migration. In previous experimental studies of grain boundary migration the driving force has been limited to a difference in stored energy across the boundary. This stored energy has been introduced into the crystal either by prior deformation1-3 or by grown-in lineage structure. A part of the energy stored in the deformed crystal is released by recovery either prior to or concurrently with grain boundary migration, thus introducing an uncertainty as to the magnitude of the driving force responsible for grain boundary migration. The grown-in lineage structure, though thermally stable during annealing, neither provides conditions under which different levels of energy may be stored in the imperfect crystal nor provides a control of orientation difference across the migrating boundary of a growing grain. Furthermore, because of variation in the lineage structure, it is difficult to determine accurately the energy stored in the imperfect crystal. Several investigations of grain boundary migration during normal grain growth have also suffered from difficulties in estimating the driving force because of uncertainties in the principal radii of curvature.~ In the present investigation the velocity of pure tilt boundaries in zone-refined aluminum bicrystals of selected orientation (40, 30, and 16 deg around the [Ill] tilt axis) has been measured in the absence of a dislocation density difference across the moving boundary, thus eliminating the previous experimental difficulties. The driving force for boundary migration is derived from a gradient of the total interfacial free energy of the migrating boundary in wedge-shaped bicrystals. A similar method was attempted by Bron and Machlin in a study of grain boundary migration in silver. However, they found that one of the crystals was deformed and consequently the motion of the boundary was partly due to a difference of stored energy across the boundary. The observed behavior of boundary velocities as affected by the driving force is examined in the light of the predictions of the current theories of grain boundary migration.7"10 The effect of boundary misorientation on velocity is compared with the theory of " which is based on a dislocation core model for high-angle boundaries. EXPERIMENTAL METHOD Seed-oriented bicrystals of zone-refined aluminum, 2.5 cm wide, 0.5 cm thick, and 12 cm long, containing tilt boundaries with a common (111) axis, were grown from the melt in the direction of this axis. Spectro-graphic analysis, reported earlier,'' indicated the purity of the crystals to be 99.999+pct. Three such bicrystals containing 16, 30, and 40 deg tilt boundaries were used. Wedge-shaped specimens were prepared from these bicrystals by spark cutting followed by electrolytic polishing. The angle of the wedge was usually 40 deg and the specimens were usually 0.25 cm thick. The intercrystalline boundary was located within 0.2 to 0.5 cm from the tip of the wedge. Fig. 1 shows a section of an oriented bicrystal containing an outline of a wedge-shaped specimen. The crystallographic directions shown in Fig. 1 represent the orientation of one of the crystals (the larger section of the bicrys-tal); the orientation of the other crystal differs only by rotation around the common [lil] axis. The parallel faces of the wedge always corresponded to the common (171) planes in both crystals, whereas the orientation of the side faces varied, depending on the misorientation angle. The bicrystal orientations were determined

Jan 1, 1970

By A. W. Bethune, L. M. Pidgeon

The equilibrium vapor pressure of zinc has been determined over the systems: ZnS(s) + Fe(s) = FeS(s) + Zn (vapor) and ZnS(s) + 2Cu(s) = Cu2S(s) + Zn (vapor) by reacting the components in an evacuated tube containing a thin copper fiber, and allowing equilibrium to be established with respect to the brass formed and the zinc in the vapor phase. The composition of the brass formed was determined and equilibrium vapor pressure values obtained from existing data. Between 850' and 1000°C, values for the first reaction ranged from 8 to 58 mm Hg; for the second reaction from 1.8 to 22 mm Hg. THE direct reduction of metallic sulphides may be indicated by an equation of the form MS + X = XS + M [1] where X is some suitable reducing agent. In the case where M is a metal of relatively high volatility, the use of vacuum will displace the equilibrium to the right at comparatively low temperatures in the face of the usually unfavorable thermochemistry. For this technique to be feasible, MS, X, and XS, must of course be nonvolatile compared to M. In this regard, zinc becomes an obvious metal both from the point of view of its volatility and the fact that most commercial ore deposits of this metal occur as the sulphide, and its direct reduction would avoid the expensive roasting process necessary prior to present day reduction methods. Two reducing agents are of interest, namely copper and iron, with the former having the added attraction of being relatively easily regenerated by oxidation in the copper converter. The possibility of these reactions has been considered before. Imbert1 patented a process whereby zinc sulphide was reduced by either copper or iron. The process as described took place at atmospheric pressure and, at the high temperatures required, the system was molten, resulting in the formation of a matte with the consequent reduction in the activities of the reactants. Peterson2 did further experimental work on these reactions and, although promising results were obtained on a small scale, attempts to increase the size of the furnace led to excessive "blue powder" formation. The early workers failed to realize the large reduction in temperature afforded by the use of vacuum and its desirable results. More recently, the direct reduction of zinc sulphide by iron has been the subject of a thermody-namic investigation by Kelley,3 and Gross and War-rington' have studied the kinetics of the reaction in the temperature range 900" to 1000°C, their experiments being conducted using vacuum and indicating that the reaction would proceed practically to completion in 1 to 3 hr, depending on the temperature. A qualitative examination of the system ZnS-Cu carried out at this University indicated that 98 pct of the theoretical zinc in the charge could be recovered in 1 hr at 1000°C when operating under reduced pressure. The purpose of this research program was to study the equilibrium of the two reactions. During the investigation, reference to the experimental work of Schenk5 was found but, due to lack of detail in the report, it was felt that it would be desirable to continue the equilibrium measurements. Selection of Method The measurement of vapor pressure of metals presents obvious problems resulting from the high temperatures which must be employed. In the present case, the production of the "reaction pressure" introduces a further difficulty since a chemical reaction must proceed before any metal vapor pressure is produced, resulting in depletion of the re-

Jan 1, 1954

By G. G. Hancock, H. H. Johnson

Hydrogen gas at atmospheric pressure is shown to cause substantial embrittlement in a martettsitic high-stretzgth steel. Subcritical crack growth is observed at very lom stresses and with high growth rates. It is suggested that an adsorbed layer of hydrogen atoms at the crack-tip surface provides the damaging hydrogen. Oxygen in small quantity, as little as O.7 pct by volume, is shown to eliminate subcritical crack growth in gas environments caused by hydrogen or water vapor. However, dissolved oxygen in water has little or no influence upon crack-growth rates. Oxygen-stopped cracks may be restarted by water, or by hydvogen or water vapor if oxygen is removed from the environment. These observations are interpreted in terms of preferential adsorption of oxygen at the crack tip and the formation of an oxide barrier. RECENT investigations1-3 have demonstrated that environmental factors are frequently responsible for delayed failure in martensitic high-strength steels. The environment causes crack initiation from preexisting cracklike flaws at unexpectedly low stresses, often far below the uniaxial yield strength. The crack then propagates at a rate controlled by the environment and the crack-tip stress-field intensity; during this period crack propagation will cease if either the stress or the aggressive environment is removed, and this may be considered a definition of subcritical crack growth. Fracture occurs when the crack has grown to a critical length, such that the fracture toughness of the steel is exceeded. Several liquids are known to cause delayed failure, but water and water vapor are the most severely damaging of the environments investigated to date.'-3 It has been suggested that hydrogen produced during corrosion reactions may be the actual embrittling agent2 The severe embrittlement caused by water vapor inevitably raises the question as to whether other gas environments are damaging. Previous work4'5 has demonstrated that high-pressure hydrogen (2250 psi or greater) is an effective embrittling medium, as measured by such criteria as reduction in area, notch tensile strength, and time to failure. Moreover, the embrittlement was eliminated by the addition of small quantities of air or oxygen to the high-pressure hydrogen, but it was not influenced by additions of nitrogen or argon.4 In the present study attention is focused upon gas environments at atmospheric pressure, in particular hydrogen and oxygen. The influence of these constituents upon the initiation and growth of a crack from a pre-existing cracklike flaw in a high-strength steel was determined directly by the electric-potential method of monitoring crack growth. MATERIALS AND PROCEDURE The H-11 steel was supplied by the U.S. Naval Research Laboratory as 3 by 12 by 0.065 in. sheet specimens with 1-in. transverse center slots. The slot tips were finished by Elox machining to a radius of 0.001 in. or less. The yield strength was 230,000 psi; chemical composition and heat-treatment schedule are given in Table I. The experimental apparatus, precracking procedure, and incremental loading sequence have been described previously.3 In the only significant change, precracking was sometimes carried out by a hydrogen embrittlement and baking technique. All experiments were conducted at room temperature; the hydrogen gas environment was purified by a commercial palladium purifier. Crack initiation and propagation were followed by the electric-potential method,'-3 using a calibration that has been recently dicussed. Stress-field intensities at the crack tip were computed by the conventional Irwin frmula; the plasticity correction was negligible and therefore not included. EXPERIMENTAL RESULTS A striking brittleness associated with molecular hydrogen at atmospheric pressure is demonstrated by the data in Table I1 and Fig. 1, which indicate that subcritical cracks in hydrogen initiate at lower stress-field intensities and propagate more rapidly than in fully humid argon. Ki and Kf are the crack-tip stress-field intensities at initiation and termination of subcritical crack growth, respectively. Termination, of course, corresponds to fracture. In both hydrogen and humid argon Ki is less than Kf

Jan 1, 1967

By D. R. Irving

Prior to this time, ample supplies of high grade mineral fillers, such as clay, have been available close to consuming centers. Now depletion of these accessible deposits, coupled with other factors such as increasing demand due to population expansion and standard of living rise, has made development of upgrading techniques attractive to producers. Various investigations being carried out by the U.S. Bureau of Mines are described. Factors making the development of improved methods for upgrading clays attractive to producers include: depletion of accessible deposits of high-grade clays, rising transportation and labor costs, more exacting specifications, and increased demand resulting from an expanding population and a constantly rising standard of living. Industry recognizes the need for upgrading presently unusable materials and has cooperated with the U.S. Bureau of Mines by providing samples from submarginal deposits and by evaluating the products that USBM upgrades. Following are some examples of recent clay beneficiation work by the Bureau. FIRE CLAY BENEFICIATION At the Rolla Metallurgical Center, Rolla, Mo., the Bureau is conducting research to develop technically and economically feasible mineral-dressing methods to remove quartz, pyrite, and other impurities from submarginal fire clays. The research has been spurred by the growing shortage of high-grade bur-ley, flint, and plastic clays in the Missouri fire-clay district resulting from increased consumption and the low rate of discovery of new deposits of adequate quality. There apparently is ample tonnage of submarginal flint and plastic clays to maintain the industry for many years if such material can be upgraded to commercial quality. Samples have been submitted to USBM by several major producers of refractories, including A. P. Green Fire Brick Co.; Harbison-Walker Refractories Co.; Mexico Refractories Div., Kaiser Aluminum and Chemical Corp.; and Wellsville Fire Brick Co. Research on some samples has been completed and still is in progress on others. The Bureau has found that specimens containing quartz or pyrite in grains coarser than the accompanying clay generally can be upgraded either with a gravity table or hydraulic cyclone. The improvement in quality is notable not only in chemical analysis but in an increase in pyrometric cone equivalent (PCE) of one or more cones. As an example, consider the results attained on a sandy flint—a fine-grained mixture of kaolin and halloysite with quartz and minor quantities of iron and manganese oxides. Chemical analysis was: 35.4 pct alumina (A12O3) and 47.8 pct silica (SiO2). The sample contained 9.2 pct free silica as quartz. The PCE was 33. The sample was crushed to -10-mesh, blunged with water, and then stage-ground to -48-mesh. The ground sample was tabled to yield a clay fraction containing 37.6 pct A12O3, and 45.3 pct SiO2 of which 2.4 pct was free silica. The PCE was 34. Recovery of alumina by this process was 73.6 pct. Concentration of the same clay in the hydraulic cyclone was even more effective in respect to quartz removal. The clay fraction analyzed 38.6 pct A12O3, 45.0 pct SiO2 of which 0.5 pct was free silica as quartz, and had a PCE of 34. Recovery of alumina was 71.9 pct. The hydraulic cyclone was also effective in reducing the pyrite content of a plastic clay which likewise was a mixture of kaolin and halloysite, plus some pyrite, quartz, and carbonaceous material. It analyzed 30.6 pct Al2O3, 52.3 pct SiO2, 2.3 pct Fe2O3, and 0.76 pct S. The sample was blunged and screened through 10-mesh to remove some coarse pyrite, then ground to -65-mesh and cycloned to remove most of the remainder. The final product recovered 81.8 pct of the alumina at a grade of 32.3 pct A12O3, 50.9 pct SiO2, 1.4 pct Fe2O3, and 0.1 pct S. The PCE was 32-1/2, compared with a PCE of 31-1/2 for the crude clay. The fine beneficiated clays, of course, must be pelletized and dried or fired to meet commercial acceptance. This operation has already been solved by industry. Little progress has been made on two remaining problems. One is that of fine-grained impurities, not separable by table or cyclone. Flotation is being investigated to solve this. Although fine pyrite is readily rejected, a satisfactory quartz-clay separation has not been attained. The second problem, a stubborn one, is that of alkali. Samples containing as much as 5 pct potas-

Jan 1, 1961

By O. Mellgren, A. M. Gaudin, P. L. De Bruyn

The adsorption density of ethyl xanthate on pyrite was determined as a function of xanthate concentration. Surface preparation of the mineral appears to have asafunctionsome effect on the subsequent adsorption process, A monolayer of xanthate on the surface is exceeded only in presence of oxygen. The effect of OH- , HS- (and x and CN- S=)and on the amount of xanthate adsorbed was investigated. Competition between OH- and X- (xanthate) ions for specific adsorption sites is indicated over a wide pH range. IN the flotation of sulfide ores, xanthates are most commonly used to prepare the surface of the mineral to be floated so that attachment to air takes place. The quantity of agent required to make the mineral hydrophobic is usually very small, of the order of 0.1 to 0.25 lb per ton of mineral. Details of the mechanism of pyrite collection are for the most part unsettled. Adsorption of collector has long been believed to involve an ion exchange mechanism as demonstrated for galena' and for chalcocite.2 In the work on chal-cocite it was also demonstrated that a film of xanthate radicals unleachable in solvents that dissolve alkali xanthates, copper xanthate, or dixanthogen was formed at the surface of the mineral. The unleachable product increased with increasing addition of xanthate up to a maximum corresponding to an oriented monolayer of xanthate radicals. Pyrite is extremely floatable with xanthate if its surface is fresh.9 ut the floatability decreases rapidly as oxide coatings increase in abundance. Pyrite shows zero contact angle when in contact with ethyl xanthate solution at pH higher than about 10.5;4 at neutrality, a contact angle of 60" is obtained at a reagent concentration of 25 mg per liter. Alkali sulfides and cyanides are pyrite depressants. In this study of pyrite collection the writers have sought to relate measured xanthate adsorption to the method used in preparing pyrite, to the presence or absence of oxygen, to concentration of hydroxyl, hydrosulfide, sulfide, and cyanide ions. The principal experimental tool has been radioanalysis," " using xanthatcx marked with sulfur 35. Experimental Materials Pyrite: Unlike most sulfides, pyrite is a poly-sulfide. The structure given by Bragg7 resembles that of sodium chloride, the iron atoms corresponding to the position of sodium and pairs of sulfur atoms corresponding to the position of chlorine. The edge of the unit cell in pyrite is 5.40 A and in halite 5.63 A. The S-S distance in pyrite is 2.10 A; the Fe-S distance, 3.50 A: and the Fe-Fe distance, 3.82 A. Natural pyrite from Park City, Utah, was used in this investigation. Pyrite 1 was obtained by hand picking pure crystals. Pyrite 2 and Pyrite 3 were obtained from finer textured crystalline material containing inclusions of silicates. The same cleaning technique was utilized for the preparation of Pyrite 2 and Pyrite 3, whereas a different cleaning technique was used for Pyrite 1. Pyrite 1 was prepared as follows: The crystals were ground in a porcelain ball mill and the 200/400 mesh fraction was separated by wet screening with distilled water, followed by washing for 1 hr with deoxygenated distilled water acidified with sulfuric acid to pH 1.5. The acid was removed by rinsing with deoxygenated distilled water on a filter until a pH of 6.0 was reached in the effluent. This filtration was carried out under nitrogen. The sample was then dried in a desiccator under nitrogen. The period of time for which this pyrite sample was in contact with water containing oxygen was about 4 hr. The specific surface as determined by the BET gas adsorption method was 582 cm2 per g. Final material assayed 53.12 pct sulfur and 46.5 pct iron (theoretical, for FeS,: S, 53.45 pct; Fe, 46.55 pct). After crushing, Pyrite 2 and Pyrite 3 were washed with 1 M HCl. rinsed, and fed to a laboratory shakinq table to remove the small amount of silicates. The concentrate obtained was ground in a laboratory steel ball mill. The 200/400 mesh fraction was separated by classification in a Richards hindered settling tube. This fraction was then given a final wash with 0.1 M HCl and deoxygenated water was filtered through the sample. The final effluent showed a conductivity equivalent to that of a solution having a salt concentration of 0.3 ppm. Aqueous hydrogen sulfide solution was then added to the sampln (about 100 ml saturated H,S solution to about 1000 g pyrite under a few hundred milliliters of water) which was stored wet under nitrogen. The sample stored in this manner showed no indication of formation of iron oxides, whereas iron oxides appeared

Jan 1, 1957

By E. W. Davis

IF the tonnage of merchantable iron ore remaining in the Lake Superior district is divided by the average of the annual shipments for the past 20 years, it will be found that this ore supply will be exhausted in approximately 35 years. This computed exhaustion period would be an indication of the future activities of the iron-mining industry if it were not for the fact that merchantable ore is being manufactured from low-grade ore, of which there is an almost inexhaustible supply in the Lake Superior district. It is true that at present only the simplest of these low-grade ores are being concentrated to any considerable extent, but progress is being made with the treatment of the more complex ores, and now, in addition to 20 washing plants on the Mesabi Range, three plants are equipped with jigs for the treatment of ore that cannot be concentrated simply by washing. The latest addition to the ore-treating plants of the Mesabi Range is a magnetic roasting and concentration plant at Cooley, Minn., in which ore that cannot be concentrated either by washing or by jigging is roasted to the magnetic state and concentrated on magnetic separators. This plant is beginning its third season of operation, and while it is too soon to make definite predictions, there are strong indica-tions that this new process may make commercially possible the manu-facture of high-grade, merchantable iron ore from large quantities of low-grade ore now considered worthless. After the iron oxide contained in an ore has been converted to mag-netite, almost any desired grade of concentrate can be produced, using magnetic-concentration equipment of the types now in general use for the concentration of the natural magnetites of New York, New Jersey and Pennsylvania. The conversion of hematite (Fe2O3) to magnetite (Fe3O4) is easily accomplished by heating the ore to a dull red tempera-ture in a reducing atmosphere. The reaction is slightly endothermic, and while it requires 200 B.t.u. to heat to 900° F. one pound of hematite ore containing 45 per cent Fe, it requires only 100 B.t.u. to convert the oxide into magnetite. In the laboratory, using small quantities of ore, this conversion is very simple, but large furnaces suitable for commercial

Jan 1, 1936

By L. M. Pidgeon, P. G. Thornhill

A LTHOUGH a considerable number of experi--ti mental investigations dealing with the roasting of sulfide minerals have been reported in the past,'"" the behavior of the single roasting particle does not appear to have received the attention it perhaps deserves. This is understandable in that the natural sulfide counterpart of the slab or sheet of metal normally used in high temperature oxidation studies is difficult to obtain and prepare, and is, in most cases, virtually impossible to maintain in one piece at roasting temperatures. The employment, on the other hand, of static aggregations of small particles in masses great enough to permit evaluation of the roasting reactions by means of thermobalances or gas analyses introduces other complications, such as localized overheating, sintering, and variable gas-solid contact. In this investigation a compromise between the above extremes was attempted. Closely sized (e.g., 30 to 40 mesh) particles of natural sulfide minerals were isothermally roasted, out of contact with one another, in an air-swept system. The progress of the roasting reactions was followed by examination in polished section of the treated particles and by X-ray and chemical analysis of their residual sulfide kernels. In this way some conclusions respecting the mechanism of roasting reactions were drawn. Roasting Furnace—The vertical tube furnace used for roasting the sulfide particles had a heated zone 10 1/2 in. long in which four retractable light gage stainless steel trays were suspended. Each tray, and an accompanying baffle, was mounted on a 2 in. lengtn of 3/8 in. stainless steel tubing. These assemblies were axially suspended on a 30 in. length of ¼ in. stainless tubing, by means of which they could be lifted from the hot zone to a water-cooled brass chamber at the top of the furnace. Since each tray unit could slide freely on the 30 in. tube, any desired number of the trays could be retained in the cooling chamber by the manipulation of a horizontal plunger during a momentary withdrawal of the four trays from the hot zone. Although three heating coils, each equipped with variable external resistance, were employed, it was found impossible to avoid a drop of 10o to 15oC between the temperature at the top tray and that of the lower three. Consequently, only the bottom three trays were used for roasting. Temperature readings inside the furnace were made by means of a thermocouple inserted down the tray support tube at tray levels, and a separate thermocouple, inserted adjacent to the middle heating coil, was connected to a Micromax temperature control unit. Experimental Method—The sulfide minerals were obtained in a form as free as possible from gangue and contaminating sulfides, and were crushed and screened to the required particle size range. The tray assembly was kept inside the furnace during the heating-up period and, when a steady temperature had been reached, the trays were withdrawn. Each tray was sprinkled with from 200 to 500 mg of the sulfide grains, and the assembly was lowered into the furnace without delay in order to minimize the time required to achieve the operating temperature. The furnace top was then clamped into position and the thermocouple inserted down the center tube to the required level. It was found that, when this procedure was used, the trays achieved their former temperature within 2 min, after which time the flow of dry air was begun at a rate of 1 liter per min, and timing of the run was started. At the required intervals of time the trays were successively withdrawn from the hot zone and stored in the cooling chamber by means of the retention plunger. On completion of the run the treated particles were removed from the trays to be set aside for examination. The roasted particles were mounted in Lucite on a Buehler press Operating at a temperature of 140°C

Jan 1, 1958

By J. H. Wernick, D. Dorsi, K. E. Benson

A LTHOUGH a considerable number of experi--ti mental investigations dealing with the roasting of sulfide minerals have been reported in the past,'"" the behavior of the single roasting particle does not appear to have received the attention it perhaps deserves. This is understandable in that the natural sulfide counterpart of the slab or sheet of metal normally used in high temperature oxidation studies is difficult to obtain and prepare, and is, in most cases, virtually impossible to maintain in one piece at roasting temperatures. The employment, on the other hand, of static aggregations of small particles in masses great enough to permit evaluation of the roasting reactions by means of thermobalances or gas analyses introduces other complications, such as localized overheating, sintering, and variable gas-solid contact. In this investigation a compromise between the above extremes was attempted. Closely sized (e.g., 30 to 40 mesh) particles of natural sulfide minerals were isothermally roasted, out of contact with one another, in an air-swept system. The progress of the roasting reactions was followed by examination in polished section of the treated particles and by X-ray and chemical analysis of their residual sulfide kernels. In this way some conclusions respecting the mechanism of roasting reactions were drawn. Roasting Furnace—The vertical tube furnace used for roasting the sulfide particles had a heated zone 10 1/2 in. long in which four retractable light gage stainless steel trays were suspended. Each tray, and an accompanying baffle, was mounted on a 2 in. iength of 3/8 in. stainless steel tubing. These assemblies were axially suspended on a 30 in. length of in. stainless tubing, by means of which they could be lifted from the hot zone to a water-cooled brass chamber at the top of the furnace. Since each tray unit could slide freely on the 30 in. tube, any desired number of the trays could be retained in the cooling chamber by the manipulation of a horizontal plunger during a momentary withdrawal of the four trays from the hot zone. Although three heating coils, each equipped with variable external resistance, were employed, it was found impossible to avoid a drop of 10O to 15OC between the temperature at the top tray and that of the lower three. Consequently, only the bottom three trays were used for roasting. Temperature readings inside the furnace were made by means of a thermocouple inserted down the tray support tube at tray levels, and a separate thermocouple, inserted adjacent to the middle heating coil, was connected to a Micromax temperature control unit. Experimental Method—The sulfide minerals were obtained in a form as free as possible from gangue and contaminating sulfides, and were crushed and screened to the required particle size range. The tray assembly was kept inside the furnace during the heating-up period and, when a steady temperature had been reached, the trays were withdrawn. Each tray was sprinkled with from 200 to 500 mg of the sulfide grains, and the assembly was lowered into the furnace without delay in order to minimize the time required to achieve the operating temperature. The furnace top was then clamped into position and the thermocouple inserted down the center tube to the required level. It was found that, when this procedure was used, the trays achieved their former temperature within 2 min, after which time the flow of dry air was begun at a rate of 1 liter per min, and timing of the run was started. At the required intervals of time the trays were successively withdrawn from the hot zone and stored in the cooling chamber by means of the retention plunger. On completion of the run the treated particles were removed from the trays to be set aside for examination. The roasted particles were mounted in Lucite on a Buehler press operating at a temperature of 146°C

Jan 1, 1958

By R. McNeill, D. E. Weiss, E. A. Swinton

In a continuous countercurrent exchange process, an alteration in any one of the operating conditions has a complex effect on the others, which can only be predicted by employing the transfer unit or the theoretical stage theory on a basis of trial and error. A simple method is described for illustrating diagrammatically the behavior of a counter-current system, the equations being simplified by means of a concept the maximum hypothetical exchange performance. An example based on a typical metallurgical system is given, in which a divalent metal is recovered from a dilute solution, the resin being regenerated continuously by a monovalent ion. Useful conclusions are drawn from a study of the theory. Practical methods for performing continuous ion exchange are discussed, and the development of equipment based on modified ore dressing jigs is described. A swinging sieve jig contactor is evaluated experimentally. DURING the last decade, the new synthetic ion exchange resins have been applied extensively in industries outside the field of water treatment, but there is no record of a continuous counter-current process operating on an industrial scale. Attempts have been made to devise a satisfactory process but many problems remain to be solved. The basic principles of continuous processes will be outlined, as well as the major problems in their operation and the progress made in the CSIRO laboratories toward the development of satisfactory industrial techniques. In the metallurgical field ion exchange resins can be used for various applications such as the recovery and concentration of valuable metals from mine waters,' the regeneration of pickling and plating liquors," the prevention of pollution by waste effluents and the recovery of the constituents from them," and the purification of valuable metals such as the rare earths by chromatographic fractionation on columns of ion exchange resins.7,8 . Turther applications undoubtedly will be found in the field of hydrometallurgy where the use of ion exchange resins would enable direct extraction of the desired metal ion from the filtered leach liquor or the leach pulp. For example, an ion exchange process has been described recently for the extraction of gold from a cyanide leach pulp." A continuous process would have advantages in many applications over the usual process employing a fixed bed and intermittent cycle. In a recovery process, it would yield a product stream of steady purity and concentration, it would waste less water in rinsing, and if the contacting apparatus were efficient less resin would be used, since each portion of the resin would be cycled as soon as it was loaded instead of lying idle until the whole bed was ready for regeneration. A very major advantage is that it would be simpler to control automatically. It is probable that continuous operation will be the key for really large scale applications of ion exchange. The flow sheet of a continuous ion exchange recovery-concentration process is illustrated diagrammatically in Fig. 1. Dilute liquor containing the valuable ion flows through the stripping section countercurrently to a moving bed of resin and leaves after a final contact with freshly regenerated resin. The resin leaves the unit almost in equilibrium with the incoming liquor and then flows to the regenerating unit where it is treated by a slow countercurrent flow of concentrated regenerant solution. The adsorbed ion is displaced from the resin and appears in the concentrated product stream. The resin then must pass through a rinse unit or section where regenerant entrained by the resin is washed back into the regeneration section by water. The regenerated and washed resin is then recycled back to the stripping section. I. Theoretical Operating Behavior of Continuous Ion Exchange Stripping System The simple theory of continuous ion exchange is analogous to that of solvent extraction and other diffusional transfer operations and is governed by the equilibrium relationship, the mass balance, the rates of mass transfer, and the contacting efficiency of the unit. Equilibrium Relationship—The relative affinity of two ions A and B, for a particular resin immersed in their solution, can be expressed by plotting compositions of the solution against compositions which exist in resin in equilibrium with those solutions, i.e. C/Co vs q/a where C, is the total normality of the solution, C is the normality of ion A in the solution, a is the total exchange capacity of the resin in gram equivalents

Jan 1, 1956

By A. H. Smallbone, R. Lathe

Considerable efforts have been expended to apply multiple regression techniques to quantitative determinations in x-ray fluorescence analysis. Multiple regression has been shown to be an effective method for correlating x-ray fluorescence intensities to elemental concentrations. Although many other techniques have been devised to solve the problem of matrix effects in x-ray fluorescence analysis, there are few as efficient in terms of storage requirement or ease of programming for digital computers as the multiple regression method. In the case of ore slurry, it has the additional advantage of allowing a unified approach to the problems of slurry density correction and matrix effect corrections. The data used in the present work were obtained with an on-stream x-ray analyzer capable of analyzing 15 slurry streams for as many as nine elements simultaneously. One of its fixed monochromators was set to accept scattered radiation at a discretely selected wavelength as a means of monitoring the slurry's pulp density. The data were reduced to obtain the correlation equations and a computer program combining slurry density with matrix corrections was written in FORTRAN IV to calculate concentrations.

Jan 8, 1969

By C. S. Smith

The solubility of tantalum at 1100°C is 0.025 pct in pure copper, 1.2 pct with 20 pct Ni, and 2.7 pct with -30 pct Ni. The solubility decreases with temperature, and the alloys are precipitation hardenable. A 79/20/1 Cu-Ni-Ta alloy reaches maximum hardness after aging at about 750°C and, if cold worked, does not recrystal-lize below that temperature. The alloys have good tensile properties at moderately elevated temperatures and, since they can be hot and cold worked nearly as easily as cupronickel, they are suggested for service at temperatures above the usual limits for copper alloys. MASSIVE tantalum dissolves extremely slowly in copper-rich alloys, and tantalum powder forms refractory surface layers which prevent its solution unless special precautions are taken. After many trials, an 80 pit recovery was consistently obtained by using a mixture of a finely divided tantalum powder with two or three times its volume of potassium tantalum fluoride (K2TaF1,-melting point about 750°C) poured in a slow stream directly on to the surface of the molten copper alloy, so that each particle would remain coated with flux and be immediately and individually wetted by the molten metal. A 50-50 nickel-tantalum alloy has a melting point of about 1400°C and small lumps of it slowly but satisfactorily dissolve in molten copper-nickel alloys as does commerical "ferro-tantalum." The high affinity of tantalum for carbon makes it necessary to avoid carbonaceous crucibles. COPPER-TANTALUM ALLOYS Castings with good shrinkage resulted when 0.1 pct Ta or more was added to copper melted under charcoal and cast in air. Copper in two-pound melts to which 0.1, 0.15, and 0.2 pct Ta had been added retained residual amounts, by analysis, of 0.025, 0.023 and 0.026 pct, respectively. The solubility of tantalum in molten copper at about 1200 °C is therefore about 0.025 pct. The electrical conductivity of these three samples was 100.48, 100.00 and 100.20 pct IACS in the annealed condition, and 98.24, 97.98 and 98.08 in the cold-drawn condition. Wires 0.080 in. in diam withstood from thirteen to nineteen reversed bends after annealing for 30 min in hydrogen at 850°C and the metal was therefore completely deoxidized. The annealing temperature corresponding to 50 pct loss of work hardness in 1 hr in a strip cold rolled 77 pct reduction was 175C, compared to 230°C for equivalent undeoxidized material. It is unusual for a decrease of recrystallization temperature to result from an alloying addition, and the tantalum probably combines with and removes some minor impurity in the original copper that restrained its recrystalliza- tion. No difficulty whatever was encountered in hot or cold rolling or drawing these ingots, and were tantalum not so expensive it would make an excellent deoxidizer for copper. COPPER-NIcKEL-TANTALUM ALLOYS Studies were made of the solubility of tantalum in a large number of copper-rich alloys, but of these significant solubilities were found only in the case of an iron alloy (which, however, separated into two immiscible liquid layers) and alloys of copper and nickel, in which tantalum is freely soluble and which were found to have interesting properties. The latter alloys are the subject of U. S. Patent No. 2,430,306 (1947). They are not at present available commercially. The Ternary Constitution Diagram—The copper-rich corner of the constitution diagram was constructed on the basis of microscopic examination of samples of various composition and treatments. The alloys were hot rolled to 0.25 in. from castings 0.625 in. thick, then annealed for 1 hr at 1000°C, quenched, cold rolled to 0.10 in. and finally annealed for various periods of time at temperatures between 800" and 1100°C in a nonoxidizing atmosphere and quenched. When the solid-solubility limit had been transgressed, particles of a compound believed to be Ni2Ta*were

Jan 1, 1960

By Frank Doxey

AT Dominion Steel & Coal Corp. it has long been recognized that continued mechanization of mine operations is necessary in the Pictou, Cumberland, and Sidney coal fields of Nova Scotia. The varied physical conditions in these fields call for special consideration of individual cases before planning is finalized. Because standard equipment cannot be procured which would operate successfully, many experiments have been necessary over the years to keep pace with the progress made in other countries. There are two mines, producing 2000 tpd, located in the Pictou coal field. The field is badly distorted and crossed by many faults. Seams are highly inclined and irregular and vary in thickness from 5 to 40 ft. Entries are difficult to maintain because of squeezing of the coal. ribs and movement of the roof and pavement. Output from the three operating mines in the Cumberland field is 3000 tpd. The field is highly inclined, inclination varying from 12° to 32°. Overlying beds consist of shales and massive sandstone lenses of extreme toughness and are responsible for bumps when the stresses are relieved by extraction. At greatest depth these are among the deepest coal workings in the world. Depth of cover ranges from 2300 to 4000 ft. This prohibits room-and-pillar working and necessitates longwall operation. Working of contiguous seams concurrently to maintain output increases the already difficult conditions. The Sydney field, with a frontage of about 30 miles, is the most important of the Nova Scotia coal fields. With the exception of one small area it is now wholly submarine. Output is approximately 21,000 tpd. Seams are 21/2 to 8 ft thick, and cover in the areas varies from 600 to 2300 ft, with an average of 1500 ft from sea bottom to the top of the seam. The seams dip in a seaward direction, pitches ranging from 6" to 4.0". The shoreline is the last place of entry to the seams and distance from the bank to the working faces is generally over 3½ miles, in some cases as much as 6½ miles. Ventilation is a problem and requires the construction of large permanent airways. Getting and Loading Coal: In 1925, in view of heavy pressures exerted by thickness of cover overlying the seams, roadways and pillars of the room-and-pillar system being worked began to break up and coal was lost. It was decided that a change in the method of extraction in areas with heavy cover was necessary, and experiments were made with many short walls and longwalls varying between 90 and 250 ft. Trial and error proved that the best operating length was between 400 and 500 ft, delivering all coal to the dip, with roadways to the face following a level course. The change-over was gradual, and the technique of roof control developed with the system, so that falls to the face are now very infrequent. An advantage of longwall mining is that it yields 95 pct extraction, especially important in coal seams of high quality or in seams where faults or disturbances restrict the workable areas. This percentage of extraction is based on the fact that the longwall advancing system takes development faces where pitch permits instead of driving headings and leaving roadway pillars. This system yields high tonnage during development and limits loss in extraction to duff left during operations. Accompanying disadvantages, on the other hand, are the heavy construction cost of main roadways and the necessity of driving all new flank face roadways through the gob. If the main roadways are driven through the solid, and large enough pillars are left on each side for protection against flank face weights, then the width of solid coal is approximately 1700 ft. This represents 10.6 pct of the coal available, or 89.4 pct extraction of the whole. These pillars, however, are of such size that they provide a useful pillar drawing area as a final operation of the mine. It may be that although the seam cannot be generally mined by the room-and-pillar method, it can be adapted to

Jan 1, 1955

By A. N. Holden

A DISLOCATION mechanism has been described by Cottrell' by which metals can yield locally, I. form Liiders bands, giving rise to a characteristic stress-strain curve with a sharp yield point and appreciable strain at constant or decreasing stress. It is undoubtedly the best mechanism that has been suggested to date." In its present development, however, the dislocation mechanism provides a more satisfying explanation for the sharp yield point than for the extensive localized flow occurring at the lower yield stress. The primary objective in this paper is to extend the dislocation mechanism to account for localized cataclysmic flow by a dislocation collision process and to give experimental evidence to support such a process. Only the yielding of iron containing carbon -will be discussed, although other metal-solute systems are known to behave similarly. Cottrell Mechanism In brief, Cottrell explains the yield point in the following way: The dislocations in iron which must propagate to produce slip usually lie at the center of local concentrations of carbon atoms, since segregation about these dislocatlons relieves some of the local stress resulting from them. A dislocation surrounded by a "cloud" of carbon atoms is thus anchored, and a higher stress is required to set it in motion than to move a free dislocation. Considering all available dislocatlons to be anchored in this fashion, the iron exhibits a yield point when the first dialocations break free and move through the lattice causing slip. This first breaking away of a dislocation enables other dislocations to break loose by "interaction" and the process becomes a cataclysm producing local deformation or Luders bands. The yield point in the stress-strain diagram for iron is absent in freshly deformed material, but returns gradually with time; the phenomenon is one aspect of what is called strain aging. The rate at which the yield point returns following straining depends on the temperature of aging. According to Cottrell the rate of return of the yield point in strained iron is limited by the rate of diffusion of carbon at the aging temperature, the mechanism is onr: of reforming the solute atmospheres around carbon-free dislocations that had stopped moving coincident with the removal of stress. If the specimen is retested immediately after straining and unloading, carbon will not have had time to diffuse to, and re-anchor, dislocations and the yield point will not occur. The carbon diffusion limitation for the rate of strain aging apparently applies if the criterion for strain aging is either the change in hardness" or the change in electrical resistance" of the strained speci- men with aging time. The possibility exists, however, that the yield point actually returns to strained iron at some rate other than that deduced from hardness or electrical resistance data. Therefore, as a preliminary experiment, the rate of yield point return in a rimmed sheet steel strained 6 pct in tension was measured at 27°, 77°, and 100°C. A plot of yield-point elongation for each of these temperatures against aging time appears in Fig. 1. The aging process is described by curves which rise to a plateau value of elongation that seems independent of temperature, but at a rate that depends on temperature. Very long times lead to a further rise in the yield-point elongation above the plateau value. However, if the later increase in yield-point elongation is ignored and the log of the time to reach half the plateau value of elongation is plotted against 1/T, a straight line results for which an activation energy of about 25 kcal pel- mol may be assigned. Within the accuracy of this sort of experiment this is approximately the activation energy for the diffusion of carbon in iron (20 kcal per mol), and the carbon diffusion limitation suggested for the yield-point return on strain aging is valid. The Cottrell mechanism thus explains in a qualitative manner the occurrence of a yield point in iron and its return with strain aging. It fails, however, to explain some of the other experimental observations that have been made of the yielding behavior of iron. For example, it is known that the yield point in iron becomes less pronounced with increasing grain size. Annealed single crystals of iron have very small yield-point elongations .if indeed they have any,' compared to a polycrystalline steel. If the only requirement for a yield point is that the dislocations in the lattice of the annealed. material be anchored by carbon atoms, the difference in the behavior of single crystals and polycrystals is not explained. That a dislocation mechanism may be entirely consistent with little or no yield point in an annealed single crystal will become apparent later when dislocation interaction is discussed. Strain aging produces a definite yield point even in single crystals. This accentuation of the yield-point phenomenon in single crystals after strain

Jan 1, 1953

By Roger V. Pierce

INTRODUCTION IN the last few years, much study has been devoted to increasing stoping efficiency. The reasons for this are shortage of manpower, shorter working hours, operating regulations, and shortages of essential materials. More efficiency in mining methods is needed to offset these adverse factors. In the past, attention has been centered upon haulage, hoisting equipment, mill flowsheets, etc., but recently the spotlight has been focused on the cost of stoping and the cost of handling materials. Scrapers and scraper hoists have thus assumed greater importance. In order to make it possible for certain ore bodies to be mined at a profit, it has been necessary to apply assembly-line practices to the streamlining of mining methods. All cycles of the mining operation must balance, and the sooner broken ore is removed from the working face, the sooner another round can be started. Since the introduction of scraper hoists in metal mines more than thirty years ago, much progress has been made in their design and application. Alert operators are now constantly watching for methods in which to apply scrapers. The saving of manpower in development work and the changing of mining methods that once employed manual handling of materials are increasing vital wartime tonnages. Essential ores from newly discovered ore bodies are moving rapidly from stopes to reduction plants-because it is now possible to develop an ore body and place it in production with a minimum of manpower and development work. Today, speed is one of the prime essentials of mining. Everywhere, there is a concentrated effort to use experienced miners in breaking rock, in timbering, and in carrying out the various phases of mining that require many months of training. Much effort is being made to mechanize the working places because experienced men naturally prefer to work where the majority of materials are handled mechanically. In line with this policy, many new men are being taught the details of scraper work, because even green men can quickly become acquainted with scrapers and thereby release older heads to perform the more skilled phases of mining. In some places it has been found economical to change from one system of stoping to another that more readily allows the application of mechanical muck-handling equipment. Where blocks are worked faster, the cost of materials, time, and labor is cut drastically. This favorable situation allows more manpower to remain on straight breaking and timbering-or straight production mining. More-over, when blocks are mined faster, more floors can often be mined vertically before the mucking or wheel floor is moved. This results in several direct savings. There are cases where it has been proved that by working a given block faster more area, longitudinally or vertically, could be mined at one time, thus permitting the reduction of development cost. (Fig. I.)

Jan 1, 1943

By K. Tangri, B. Swaroop

Simultaneous measurements of lattice (elastic) strain by X-ray line shift method and total strain with an electrical strain gage have been carried out on polycrystalli,ne nickel with the help of a specially designed tensometer attachment for the X-ray dif-fractometer. During the initial stages of deformation, the rate of increase in lattice strain closely follows the total strain until the plastic strain sets in. From then onwards, the two strains deviate from each other and with further increase in afiplied stress the rate of increase of lattice strain eventually decreases. Depending upon the mode of unloading, both compressive and tensile strains have been observed in nickel deformed up to 0.29 pct strain. These results have been explained on the basis of the effects of clustering of dislocations and also the production and behavior of point defects during loading and unloading, respectively. POLYCRYSTALS deformed plastically in a uniaxial tension test show residual lattice strains (hereafter referred to as RLS), which broaden the X-ray line profiles and shift their peak positions.''4 It is known that uniform straining of a crystal lattice (macrostrain) produced movement of ddfraction line peaks, whereas nonuniform straining (microstrain) causes line broadening.= Though the nature and origin of RLS in deformed metals is not yet clearly understood, these are believed to be due to the presence of some form of a locked-up stress system. Several possible detailed interpretations6'" of the stress system have been proposed by various workers and common to all these interpretations is the assumption that different parts of the aggregate have different tensile yield stresses; e.g., that a part A yields under a lower applied stress than a part B. Therefore, during the deformation process, the elastic strain in A will be less than that in B, and, after completion of deformation, B will tend to contract further than A but will be prevented from doing so by the restraining influence of A. Thus, when equilibrium is reached, A will be in compression while B is in tension. Although there is general agreement on the correctness of this argument, controversy still exists as to the exact nature of the parts A and B in a deformed metal. In an alternative approach to the creation of parts A and B, many other investigators have assumed that the RLS observed in unloaded specimens are in some way connected with the lack of proportionality between lattice strain and applied stress in the region above the yield stress." cullity13 has prepared a schematic summary of the results of previous workers, mainly Smith and Wood,2'11 which suggests that, above the elastic limit, the lattice strain may increase less rapidly with respect to strain, or may even decrease with increase in applied stress. If the applied stress is then decreased, the lattice strain would decrease along a line parallel to the loading line in the elastic region and thus produce a compressive residual lattice strain after unloading. At equilibrium, to balance these compressive stresses? there would have to be regions under tensile stress which may well be the grain boundaries or the substructure walls formed during deformation. The present investigation was undertaken to gain a better understanding of the nature and origin of RLS and also to experimentally verify the salient features of this hypothesis. EXPERIMENTAL Stress and strain determinations were made with a specially constructed tensometer attachment for the X-ray diffractometer, Fig. 1: which permits the following measurements simultaneously on a sheet specimen under uniaxial tension at various stress levels during loading and unloading: a) lattice strains, E=. in a direction perpendicular to the direction of pulling (x direction) from shifts in X-ray line peak positions; b) the total strain, ex, in the direction of pulling. with the help of an electrical strain gage affixed to the back of the specimen directly below the area irradiated by X-rays: and c) the applied stress, with the help of a load cell which consisted of a calibrated stainless-steel sample of dimensions identical to those of the specimen under investigation, and to which it was coupled. Because of its high stacking fault energy.I4 nickel

Jan 1, 1970

By Anton Sawatzky, Erich Vogt

By means of mathematical solutions to the appropriate diffusion equations, we describe the kinetics of the thermal diffusion of hydrogen in Zircaloy-2 for the various temperatures and concentrations encountered in a heavy water moderated reactor. When the hydrogen concentration is below terminal solid solubility only the a Phase is present. Redistributions are then described in terms of the characteristic functions of the difhsion equation. For higher concentrations both the a and 6 phases are present. We assume the two phases to be always in equilibrium. For moderately small hydrogen concentrations exact solutions of the two-phase equation approach the the approximate solutions derived by Sawatzky and for all concentrations the exact solutions exhibit the qualitative features of his result: the two-phase concentration increases with time, everywhere; in the absence of a hydrogen current at the hot end of the sample an a-phase region always exists there; the interface of the a + 6 , a-phase boundary moves toward the cold md of the sample and the hydrogen concentration is discontinuous at the interface. Simultaneous solulions of the a and a + 6 hydrogen distributims and of the concomitant interface motion are obtained and compared to the observations of Sawatzky and Markowitz. The kinetics of the hydrogen diffusion process are shown to lead to m apparent heat of transport of the a phase which is lower than the actual value (even for samples with long anneals) thus resolving at least partially the disparity between experimental measurements of this quantity. A number of recent papers1"4 have reported measurements on the diffusion and redistribution of hydrogen in Zircaloy-2 under temperature and concentration gradients. These studies were instigated by problems arising from increasing use of Zircaloy-2 as a fuel element cladding material in pressurized-water power-reactors. The Zircaloy-2 picks up hydrogen during the operation of the reactor: the consequent precipitation of zirconium hydride in the Zircaloy-2 has pronounced effects on its mechanical properties. The Purpose of the present Paper is to describe the kinetics of the thermal diffusion of hydrogen in Zircaloy-2 for the various hydrogen concentrations and temperatures likely to be encountered in reactors. When the hydrogen in the Zircaloy-2 is entirely in the solid solution phase (a phase), the differential equation for thermal diffusion is well known and the redistribution can be described by standard mathematical methods some of which are given in Section 11 below. The treatment of the a + 6 (hydride) region differs from the earlier treatment of sawatzky3 by taking into account several modifications of the two-phase diffusion equations as suggested to us by Markowitz4 and Kidson.5 Even with the modifications the results obtained in our more complete treatment are substantially the same as those found by Sawatzky. As we show, the difference between the results of sawatzky3 and kIarkowitz4 is largely due to the difference in the geometry of their experiments. In the next section we derive the modified two-phase equation for arbitrary geometry. The exact solution of this equation is given in Section III for linear and cylindrical geometry. It is shown that Sawatzky's approximate solution is quite accurate for almost all the temperatures and hydrogen concentrations which are actually encountered. In Sawatzky's approximate theory the hydrogen concentration everywhere in the two-phase region increases continuously. As his paper pointed out, this result, together with hydrogen conservation, implies that in a sample with no hydrogen current flowing into the hot end a two-phase region is always accompanied by a single-phase region at the hot end of the sample. The net hydrogen gain in the two-phase region is supplied by the decrease of hydrogen in the single-phase region and by the movement of the (a, a + 6) boundary toward the cold end of the sample. Hydrogen conservation at the (a, a + 6) boundary leads to a discontinuity in the concentration and its derivative there. In Section IV it is shown how these qualitative features of the thermal diffusion kinetics arise from the simultaneous solution of the a-phase differential equation and a + 6 phase equation. Methods derived by us previously for the solution of the redistribution in both phases and the accompanying motion of the boundary are used to obtain approximate solutions valid for large times. The solutions account well for the observed redistributions of Sawatzky and Markowitz. It is shown how the continuing motion of the (a, a + 6) boundary leads to measured values of the heat of transport lower than the actual value, even for specimens with relatively long anneals.

Jan 1, 1963

By R. E. Cook

This paper presents results of a study to determine to what extent errors in estimated free gas saturation affect the results of static pressure calculations from build-up curves in two-phase systems. Use is made of the method of pressure build-up curve analysis developed by Miller, Dyes and Hutchinson1 , with modifications for two-phase flow advanced by Perrine2. Cornhined fluid compressibility of gas-oil mixtures is included, with special attention given to the discontinuity that occurs in fluid compressibility at the saturation pressure and its effect on static pressure calculations from build-up curves. The sensitivity of static pressure calculations to errors in gas saturation is considered over wide ranges of formation and fluid properties and reservoir pressure conditions. It is concluded from this study that for analyses of two-phase pressure build-up curves (I) the use of under-saturated crude compressibility data, even when producing GOR's are at or near solution values, will result in very significant negative errors in calculated static pressrrre, and (2) except for large values of buildup curve .slope and/or very low crude gravities, rather large errors in calculated gas saturation can he tolerated without .significant effect on calculated static pressures. The conc1usions arc qualified as referring only to the effects of gas saturation and excludc possible errors in the nuumerous other factors that enter into the ancllysis of pressure build-up curves. INTRODUCTION As part of the analysis of basic data for several major Eocene reservoirs in the Bolivar Coastal field (Western Venezuela), it was necessary to analyze a large number of pressure build-up curves for oil permeability and static reservoir pressure. From performance data it became apparent that gravity segregation is very active in these reservoirs. As a result of this performance it was recognized that the free gas saturation within the oil leg of the reservoirs would be small but would be subject to certain errors in calculation due to the inaccuracy of relative permeability data in the low gas saturation region. Since all of the various methods of build-up curve analysis were originally derived for sin- gle-phase, constant compressibility systems, and since the presence of a small but rather indeterminate gas saturation would greatly affect the reservoir fluid compressibility, a study was undertaken to investigate the effect of gas saturation on the results of two-phase pressure build-up analyses. TWO-PHASE BUILD-UP CURVE ANALYSIS The method of analysis used in this investigation is basically that originally developed by Miller, Dyes and Hutchinson1, with modifications for two-phase flow advanced by Perrine2. Miller, et al, showed that the original equation for single-phase permeability away from the well bore, held for two-phase flow except that the calculated permeability is the effective permeability to oil away from the wellbore. This finding was acknowledged by Per-rine, who also presented the equation for effective gas permeability* as Since Eqs. 1 and 2 contain no terms influenced by fluid compressibility, no distinct problem should develop in calculating permeability from two-phase build-up curves. In regard to the calculation of static reservoir pressure from build-up curves in single-phase systems, the modified Miller, et al, method as presented by Perrine is briefly as follows. 1. From build-up curve data, PVT data for the reservoir fluid and the well spacing and sand thickness, calculate the climensionless shut-in time corresponding to the highest value of actual shut-in time (t, hrs) on the straight-line portion of a semi-log build-up curve. Dimenrsionless shut-in time is related to actual shut-in time by the relation, t- 0.000264 kt 2. From a graph of dimension less shut-in time vs dimensionless pressure difference (Fig. 1 1 ) find the value of dimensionless pressure difference corresponding to the value of 7 calculated under (1) above. Dimen-

By Michael Tenenbaum

THE importance of semikilled steel as a high tonnage grade has long been recognized. The increasing severity of the applications for which semikilled steel is used makes it desirable to obtain further information regarding the features of open hearth practice and ingot structure that can affect steel performance in subsequent rolling and fabricating operations. Accordingly, an investigation is being carried out studying the structure of various types of semikilled steel ingots. This paper reviews some of the observations made and information gathered in this investigation. By definition, semikilled steel may be considered to include all nonrimming steels in which the natural shrinkage occurring during ingot solidification is offset to an appreciable degree by gas formation. Accordingly, this type of steel includes the entire range intermediate between rimmed and fully killed steel. The resultant ingot structures range from those of steels that have been capped to suppress a weak rimming action to those steels that have been killed in the furnace or ladle to the extent that no mold deoxidizer is required. A survey reported to the 1946 AIME Open Hearth Conference' reviewed the practices being used at 28 major plants for the deoxidation of semikilled steel. It was found that most plants divide the deoxidizing additions between the ladle and the mold many of the plants adding only minor amounts of deoxidizer to the bath or none at all. Silicon, aluminum and titanium were the only elements used for ladle and mold deoxidation. Mold deoxidizers were added as capping additions or as uniformly fed additions. Several practices were reported in which the steel was deoxidized in the ladle to the extent that little or no mold deoxidation was needed. No reference was made to grades capped by using special mold designs or special closures over the top of molds. Accordingly, specific examples of this type of practice will not be considered in this paper. It appeared from this survey of deoxidation practice that semikilled steel can be divided into three types, namely, capped semikilled, in which aluminum is added near the end of pour or after shut-off, intermediate semikilled steels, in which aluminum is fed uniformly throughout the pour, and semikilled steels requiring no mold deoxidation. In an attempt to cover the ingot structures obtained by the range of practices reported in the preceding survey, three series of experimental heats were made using increasing amounts of aluminum, silicon and titanium for ladle deoxidizers. The ladle deoxidizers were added as stick aluminum, 50 pct ferrosilicon and 20 pct ferrotitanium (4 pct carbon), respectively. Two methods of studying ingot structures were used. Ingots were removed from production, burned longitudinally and machined to allow observation of the central pipe and coarse blowholes. During the machining operation much of the finely porous subsurface structure was obliterated. Accordingly, observations were also made

Jan 1, 1947

By H. D. Hodges, J. K. Godbey

In order to better understand the fracturing process, bottom-hole pressures were measured during a number of typical fracturing operations. A recently developed system was used that allows simultaneous surface recording of both the bottom-hole and wellhead pressures on the same chart. The results from six fracruring treatments are summarized on the basis of the pressure data obtained. Al-though no complete analysis is attempted, the value of accurate pressure measurements is emphasized. Important characteristics of the bottom-hole pressure record do not appear at the wellhead because of the damping effect of the fluid-filled column. In four of the six treatments described, the formations apparently fractured during the initial surge of pressure with only crude oil in the well. The properties of the fluids used during the treatments are given and the fluid friction losses are obtained directly from the pressure records. This technique is also shown to be adequate for determining when various fluids, used during the process, enter the formation. INTRODUCTION Hydraulic fracturing for the purpose of increasing well productivity is now accepted in many areas as a regular completion and workover practice. Numerous articles have appeared in the literature discussing the various techniques and theories of hydraulic fracturing'. In general, three basic types of formation fractures are recognized today. These are the horizontal fracture, the vertical fracture, and fractures along natural planes of weakness in the formation'. Any one or all three of these fracture types may be present in a fracturing operation. However, with only the wellhead pressure record as a guide, it is difficult at best to determine if the formation actually fractured, and is almost impossible to determine the type of fracture induced. These difficulties arise in part because the wellhead pressure record, especially when fracturing through tubing, does not accurately reflect the pressure variations occurring at the formation. Several factors contribute to this effect and preclude the possibility of using the wellhead pressure as a basis for accurately calculating the bottom-hole pressure. These factors are: 1. The compressibilities of the fluids which damp the pressure variations. 2. The changes in the densities of the fluids or apparent densities of the sand-laden fluids. 3. The flowing friction of the various fluids and mixtures, which is dependent on the flow rates and the condition of the tubing, casing, or wellbore. 4. The non-Newtonian characteristics of a sand-oil mixture and its dependence upon the fluid properties, the concentration of sand, and the mesh size used. 5. The unknown and variable temperatures throughout the fluid column. Because of these reasons it was determined that in order to obtain a more accurate knowledge of the nature of fracturing, the bottom-hole pressure must be measured along with the pressure at the surface during a fracturing treatment. Even with accurate pressure data, a reliable estimate of the nature of fracturing is still dependent upon knowledge of the tectonic conditions. However, the hydraulic pressure on the formation is basic to any approach to a complete analysis. In order to accomplish this objective a system was developed to record the wellhead and bottom-hole pressures simultaneously at the surface. By recording both pressures on a dual pen strip-chart recorder, it was possible to greatly expand the time scale so that rapid pressure variations would be faithfully recorded. By such simultaneous recording, time discrepancies inherent in separate records are eliminated, thus overcoming one of the most difficult problems associated with bottom-hole recording systems. This paper illustrates the results obtained by using this system during six typical fracturing operations. All of these tests were taken in wells that were treated through tubing. By a direct comparison of the wellhead and bottom-hole pressures, the importance of obtaining complete pressure information during a fracturing treatment is emphasized. THE INSTRUMENTATION AND PROCEDURES The bottom-hole pressure measuring instrument consisted of a pressure-sensing element, a telemetering section, and a lead-filled weight or sinker bar. The pressure-sensing element used was an isoelastic Amerada pressure-gauge element. By using an isoelastic element, no temperature compensation was necessary in the tests described, since the temperature was believed to be well below the maximum temperature limit of 270°F. The rotary output shaft of this helical Bourdon tube element was coupled to a precision miniature potentiometer. The rotation of the pressure-gauge shaft thus changed the resistance presented by the potentiometer