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By J. B. Clark, A. J. McEvily, R. L. Snyder

The nonuniform distribution of precipitate particles has been recognized as a leading factor contributing to the relatively low fatigue resistance of aluminum alloys. The structure of many of these alloys is characterized by narrow precipitate-free zones adjacent to the grain boundaries. Alloys with such zones exhibit a tendency for brittle inter crystalline fracture. The interrelation between this type of structure and mechanical properties was investigated in an Al-10 wt pct Mg alloy. It was found that deformation during fatigue occurs preferentially along these zones and cracks initiate there. In Al-10wt pct Mg, the zones were found to be supersaturated even after extensive general precipitation and are due to the absence of proper precipitate nuclei in the region near the grain boundaries. Cold working the alloy prior to aging improves the mechanical properties by inducing precipitation within the zones and also by jogging of grain boundaries. The mode of fracture is changed from brittle inter crystalline to more ductile trans granular fracture. THE process of fatigue is highly structure sensitive, with the strength of the whole often dependent upon some localized discontinuity, either geometrical or metallurgical in nature. Much has been learned about the role of geometrical discontinuities, e.g., notches, in fatigue, but with the exception of the effects of inclusions or the shapes of carbides, relatively little is known about the specific effects of discontinuities in metallurgical structure such as nonuniform precipitation. In most age-hardening aluminum alloys, metallo-graphic studies have shown that the extent of precipitation adjacent to grain boundaries is much less than that which occurs in the interior of the grains. The width of these almost precipitate-free regions, which are sometimes called denuded zones, and the extent of solute depletion within them, are dependent upon the particular alloy and its aging treatment. It has been observed1 that these zones are relatively soft with the result that plastic deformation takes place preferentially within them. It has also been shown 2-4 that there exists a tendency for intercrys- talline cracking in fatigue when such zones are present. It is of interest to note that Broom et al.2,3 were able to reduce the incidence of this type of failure in an A1-4 wt pct Cu alloy by stretching the material 10 pct prior to aging. In the present study, the effects of precipitate-free regions on the fatigue properties of an A1-10 wt pct Mg alloy were studied in detail, and the effects of deformation prior to aging on the nature of the precipitation process as well as on fatigue properties were also investigated. MATERIAL AND PROCESSING An A1-10 wt pct Mg alloy was selected for this study, because it was known that well-defined precipitate-free regions along the grain boundaries are readily obtained in this alloy after aging at 200oC.5 The starting materials were 99.998 pct A1 and singly sublimed magnesium of about 99.9 pct purity. The aluminum was induction melted in a graphite crucible, and then the magnesium addition was immersed until dissolved. Chlorine gas was then bubbled through the molten alloy for 4 min to degas the melt, after which the melt was cast at a pouring temperature of 730" to 760°C into a cold, graphite-coated, tapered steel mold. Since A1-Mg alloys are difficult to homogenize,5 special care was taken to obtain a uniform composition. Two-in. cubes were cut from the ingot and heated at 446°C for 30 min. These cubes were then hot forged approximately 35 pct in each of the three cube directions and homogenized for 16 hr at 446°C. Sheet specimens were then obtained by pressing 40 pct and rolling 35 pct per pass with reheating between reduction steps to a final thickness of approximately 0.10 in. The sheet was then solution treated for 16 hr at 446°C and water quenched. The age hardening behavior of this material at 200°C was then determined, and the results are shown in Fig. 1. The age hardening of this alloy when subjected to cold work prior to aging is also shown in this figure. Preliminary work indicated that extensive deformation after quenching was required to affect drastically the precipitate-free regions in this alloy, and a rolling reduction of 50 pct was chosen. For purposes of comparison the following three conditions were studied: a) Solution treated, quenched, and aged 20 hr at 200°C

Jan 1, 1963

By R. R. Estill

THE rank of the Ruhr coal ranges from a high volatile bituminous coal to an anthracite, depending to some extent on the original depth of the seam. The average Ruhr coal corresponds to a soft bituminous American coal of a coking quality. The average thicknesses of individual coal seams being mined are also comparable (59 in. against 65 in. in the United States). However, consideration of seam conditions and mining conditions other than those just mentioned emphasizes differences rather than similarities with United States soft coal. In general, the Ruhr seams now being mined are much more folded and inclined than American seams. Dips of 20' and 30" are common in seams now being worked, and 30 pct of the coal reserves in the district are in seams dipping more than 35". Only on the tops and bottoms of folds do we find rather flat coal seams. In addition to the folding there is extensive displacement by cross faulting plus a certain amount of strike faulting of an overthrust nature, which results locally in doubling or omission of seams. Because of the long history of mining in the Ruhr, nearly all coal lying near the surface has long since been mined out, and we find that the average depth of mining is at present about 2300 ft below the surface. Deep mining, folding, and faulting result in seam conditions requiring a great deal more roof support than one finds in American soft coal mines. In fact only in the anthracite district and the Rocky Mountain and Pacific coal fields do we find somewhat similar conditions. It is easy to say, therefore, that the problem of mechanization of coal cutting and loading in the German mines is quite different from that which we have so effectively met in America with our mobile cutters and loaders, duck bill loaders, and a room and pillar system of mining our drift and slope mines. Partly because of more limited coal reserves, the traditional German mining system is largely the longwall method, which gives an almost complete coal recovery. Backfilling must be extensively practiced to protect the longwall faces, the over and underlying seams and workings, and especially the surface industrialized areas and barge canals. The German engineers have accordingly concentrated their efforts on the design of cutters, loaders, and conveyors suitable to longwall methods rather than room and pillar methods. Undercutters with cutter bars like American models have been in use in the Ruhr since well before World War 11. In 1941 they accounted for 8.5 pct of the production. This percentage, of course, includes coal which was undercut but nevertheless had to be broken down with air hammers or with explosives. The most common of these cutters is the Eickhoff Standard cutter (see fig. 1). This machine does about 95 pct of the undercutting in the Ruhr today, and is available with either compressed air or electrical power and in at least four different sizes. A variation of the cutter is this one with two cutter bars (fig. 2). At the end of 1947 about 200 of these machines and similar cutters were accounting for 13.2 pct of the total production, a production which was, however, only 60 pct of the 1941 production rate, so that the actual cutter tonnage was only up to a small amount over 1941. In 1941 about 3 pct of the production was accounted for by shearing machines making their cut perpendicular to the longwall face. They were similar to those used in the States. These machines are today considered obsolete and now account for only 0.7 pct of the total production. They are located at only a few mines and at present do not seem to have much of a future in the Ruhr. For the future, the Ruhr miner is looking forward to rather extensive mechanization of face work, with two major types of equipment being developed almost simultaneously. On one hand there is the development of cutter loaders for use in relatively hard coal. They represent the further extension of ideas developed after relatively long experience with the Eickhoff cutter. On the other hand there has been since 1942 an intense interest in the Ruhr in the development of face-stripping methods, particularly by the Kohlenhobel (coal plow) and its modification. At the end of 1947 these cutter loaders, Kohlen-hobels and scrapers together were actually accounting for only about 1.4 pct of total production while air hammers still broke 77.1 pct and as much as 1.2 pct was actually broken by hand picks. However,

Jan 1, 1951

By F. H. Buttner, H. Udin, J. Wulff

Using a modified Udin, Shaler, and Wulff technique, the surface tension of gold Udin, purified helium was found to be 1400 ± 65 dynes per cm for the temperature range 1017° to 1042°C. IN the original Udin, Shaler, and Wulff technique for measuring the surface tension of copper: variously weighted wires were allowed to extend or contract in a copper cell held at elevated temperatures in vacuum. By plotting stress vs. strain for a wire array in one test, the stress at zero strain is obtained. This is the point where the contractile forces resulting from surface tension are balanced by the applied load, according to the expression: y = e=o r [1] where y is the surface tension in dynes per cm; a,,,, the stress at zero strain in dynes per cm; and T, the radius of the wire in cm. The assumption that the wires deform viscously permits the drawing of a straight line through the points on the stress-strain plot. Justification of the assumption has received further experimental support recently.'-' The presence of grain boundaries in the wires requires a correction to the original expression used." Thus: y = d4=T [l- (dl) (ar)Y1 [2] where, n/l is the number of grain boundaries per unit length, and a, the ratio of grain boundary tension to free surface tension. Alexander, Kuczynski, and Dawson in studying the creep of gold wire in vacuum were unable to obtain reproducible values of the surface tension of gold. In plotting stress vs. strain for progressively longer times, they found that the stress at zero strain drifted with time from positive stress values to negative values. Similarly, for the surface tension of silver, reproducible values were obtained only when a purified helium atmosphere was substituted.' Evidently the evaporation rate of silver in vacuum is too high at the temperatures employed to obtain solid-gas equilibrium even in a similar metal enclosure. Thus reproducibility of results is lost. Experimental Procedure The experimental procedure was much the same as that originally developed by Udin, Shaler, and Wulff with a few modifications and improvements. For greater accuracy in strain measurements, knots gave way to cut gage marks as shown in Fig. 1. These were made with a hand-driven lathe in which razor blades serv'ed as cutting tools. Also a more precise cathetometer with a screw accurate to 0.00015 cm was used. The tests were conducted in an atmosphere of purified .helium rather than in vacuum in order to avoid possible evaporation difficulties. Five mil wire of high purity gold (99.98 pct) was used. After cutting in the gage marks, each wire of a series of about 12 was differently loaded by welding a gold ball to one end. This was done by dipping the end of the wire in a cooling gold droplet, previously melted on a charcoal block with a No. 2 acetylene torch. The other end of the wire was strung through a hole in a gold lid and twisted over the edge to hold the wires fixed and in suspension from the lid. The lid and mounted wires were then dipped in pure ethyl alcohol to dissolve any skin oils and dirt on the surface of the wires due to handling. Finally the lid was put in place on an alundum crucible lined with gold so that the wires hung freely within the gold-lined chamber. This whole assembly was next heated in a quartz nichrome wound tube furnace and heated for a few minutes at 600°C to soften the wires. After this anneal the wires were easily straightened with tweezers. The wire assembly was finally annealed 10" to 25°C above the subsequent test temperature for 2 hr. This treatment allowed the grains to grow to equilibrium size and shape. After the anneal, the lid was mounted in front of the cathetometer. The gage length was measured by sighting the 40 power microscope on the upper lip of the lower gage mark for the first reading, then traveling up to the lower lip of the upper gage mark for the final reading. This procedure was repeated four times to give an average gage length value. In this manner the annealed gage length and the final gage length could be measured to determine the strains. During all measurements, grain counts were made.

Jan 1, 1952

By R. V. Mrazek, F. E. Block, S. B. Knapp

The mixed homogeneous and heterogeneous kinetics of the thermal decomposition of tungsten hexacarbonyl were studied by employing a batch reactor. The system was such that a sample of tungsten hexacarbonyl could be injected into the preheated reactor, and the progress of the reaction followed by a simple pressure measurement. Both the homogeneous and heterogeneous reactions were found to be first order, and approximate activation energies were determined for each reaction. It is shown that the dis-proportionation of carbon monoxide to give carbon and carbon dioxide cannot be the source of carbon in tungsten deposits prepared by this reaction. The kinetics of the thermal decomposition of tungsten hexacarbonyl have been investigated as part of a continuing study by the U.S. Bureau of Mines on the decomposition of organometallic compounds. Reactions involving the thermal decomposition of metal carbonyls have a potential application in the preparation of pure metals and fine metal powders. Indeed, it was these applications which provided the impetus for much of the early work involving the carbonyls of nickel1 and iron.' The relative lack of study of other metal carbonyls can be traced to the comparative difficulty in synthesizing these compounds. The most common use for tungsten hexacarbonyl has been as an intermediate in vapor-phase plating.7'8 However, attempts to obtain a carbon-free deposit of tungsten by this method have not been successful, and some investigators have taken advantage of the carbon contamination and used this process to form tungsten carbide deposits.lo Other investigators have studied the thermodynamic properties11"14 and molecular structure of tungsten hexacarbonyl. However, very little is known about the kinetics of this thermal decomposition, the mechanisms involved," or the source of carbon in the resulting plate. In contrast, studies have been made of the kinetics of the thermal decomposition of nickel tetracarbonyl, iron pentacarbonyl, and molybdenum hexacarbonyl.'l It has been found that these thermal decompositions occur by a mechanism which is partially heterogeneous in nature. Information available on the equilibrium constants for the decomposition of tungsten hexacarbonyl was used to determine a temperature range, 500" to 560°K, in which the reaction could be expected to be essentially complete. APPARATUS The apparatus used allowed the injection of a sample of tungsten hexacarbonyl into a preheated batch reactor and the use of a simple pressure measurement to follow the progress of the reaction in the sealed reactor. The pressure was sensed by means of a pressure transducer (Consolidated Electrodynamics Corp., 0.3 pct)* capable of operating at the *Reference to specific products is made to facilitate understanding and does not imply endorsement of such brands by the Bureau of Mines._______ reaction temperature. This type of sensing element was chosen to avoid the problem of condensation of the sublimed carbonyl in the capillary tubing leading to any type of remote pressure-sensing device. stirring was provided by rotating the entire apparatus. Glass beads placed in the reactor provided a pulsating agitation. To minimize thermal gradients in the reactor walls, the reactor was constructed of aluminum. The support tube which held the reactor in the furnace was thin-walled stainless steel to minimize heat conduction out of the reactor. As a result of these measures, a nearly uniform temperature (°C) was maintained throughout the reactor. Fig. 1 is a schematic diagram of the apparatus. The small gear motor rotated the entire apparatus at about 200 rpm. The bearings shown at the ends of the air cylinder were perforated to allow air to be fed to the charging piston and to allow inert gas to be fed to the reactor during the preheating period. The sample was simultaneously injected and sealed inside the reactor by operation of the air piston. Fig. 2 shows a cross section of the air cylinder and the adjoining portion of the support tube leading to the reactor. The sample carrier is shown in place at the right-hand end of the injection rod extending from the air piston. The piston is shown in the retracted position, as it would be prior to the start of an experiment. The small Teflon gasket which encircled the sample carrier at the end of the injection rod sealed the reactor when the sample was injected. This seal was maintained throughout the test by maintaining air pressure on the piston. The sample carrier was a 2-in. section of thin-walled, -in.-diam nickel tubing with an internal blank about 1 in. from the base and with the base end sealed.

Jan 1, 1969

By W. M. McKewan

Magnetite pellets were reduced in flowing hydrogen at pressures up to 40 atm over a temperature range of 350° to 500°C. The rate of weight loss of oxygen per unit area of the reaction surface was found to be constant with time at each temperature and pressure. The reaction rate was found to be directly proportional to hydrogen pressure up to 1 atm and to approach a maximum rate at high pressures. The results can be explained by considering the reaction surface to be sparsely occupied by adsorbed hydrogen at low pressures and saturated at high pressures. PREVIOUS investigation1,2 have shown that the reduction of iron oxides in hydrogen is controlled at the reaction interface. Under fixed conditions of temperature, hydrogen pressure, and gas composition, the reduction rate is constant with time, per unit surface area of residual oxide, and is directly proportional to the hydrogen pressure up to one atmosphere. The reduction rate of a sphere of iron oxide can be described3 by the following equation which takes into account the changing reaction surface area: where ro and do are the initial radius and density of the sphere; t is time; R is the fractional reduction; and R, is the reduction rate constant with units mass per area per time. The quantityis actually the fractional thickness of the reduced layer in terms of fractional reduction R. It was found in a previous investigation2 of the reduction of magnetite pellets in H2-H,O-N, mixtures, that the reaction rate was directly proportional to the hydrogen partial pressure up to 1 atm at a constant ratio of water vapor to hydrogen. Water vapor poisoned the oxide surface by an oxidizing reaction and markedly slowed the reduction. The enthalpy of activation was found to be + 13,600 cal per mole. It was also found that the magnetite reduced to meta-stable wüstite before proceeding to iron metal. The following equation was derived from absolute reaction-rate theory4,8 to expfain the experimental data: where Ro is the reduction rate in mg cm-2 min-'; KO contains the conversion units; Ph2 and PH2O are the hydrogen and water vapor partial pressures in atmospheres; Ke is the equilibrium constant for the Fe,O,/FeO equilibrium; Kp is the equilibrium constant for the poisoning reaction of water vapor; L is the total number of active sites; k and h are Boltzmann's and Planck's constants; and AF is the free energy of activation. Tenenbaum zind Joseph5 studied the reduction of iron ore by hydrogen at pressures over 1 atm. They showed that increasing the hydrogen pressure materially increased the rate of reduction. This is in accordance with the work of Diepschlag,6 who found that the rate of reduction of iron ores by either carbon monoxide or hydrogen was much greater at higher pressures. He used pressures as high as 7 atm. In order to further understand the mechanism of the reduction of iron oxide by hydrogen it was decided to study the effect of increasing the hydrogen pressure on rebduction rates of magnetite pellets. EXPERIMENTAL PROCEDURE The dense magnetite pellets used in these experiments were made in the following manner. Reagent-grade ferric oxide was moistened with water and hand-rolled into spherical pellets. The pellets were heated slowly to 550°C in an atmosphere of 10 pct H2-90 pct CO, and held for 1 hr. They were then heated slowly to 1370°C in an atmosphere of 2 pct H2-98 pct CO, then cooled slowly in the same atmosphere. The sintered pellets were crystalline magnetite with an apparent density of about 4.9 gm per cm3. They were about 0.9 cm in diam. The porosity of the pellets, which was discontinuous in nature, was akrout 6 pct. The pellets were suspended from a quartz spring balance in a vertical tube furnace. The equipment is shown in Fig. 1. Essentially the furnace consists of a 12-in. OD stainless steel outer shell and a 3-in. ID inconel inner shell. The kanthal wound 22 in. long, 1 1/2, in. ID alumina reaction tube is inside the inconel inner shell. Prepurified hydrogen sweeps the reaction tube to remove the water vapor formed during the reaction. The hydrogen is static in the rest of the furnace. The sample is placed at the bottom of the furnace in a nickel wire mesh basket suspended by nickel wire from the quartz spring. The furnace is then sealed, evacuated, and refilled with argon several times to remove all traces of oxygen. It is then evacuated, filled with

Jan 1, 1962

By J. F. Martin, L. M. Melnick, R. Rapp, R. C. Takacs

Recent studies on the determination of hydrogen in steel have shown that the hot-extraction method for removing hydrogen from a solid sample is preferable to its removal from a molten sample by vacuum fusion or by fusion in vacuum with tin. A number of techniques are available, however, for determining the hydrogen so extracted. They include: thermal conductivity, gas chromatography, pressure measurement before and after catalytic oxidation of the hydrogen to water and removal of the water, and pressure measurement before and after diffusion of the hydrogen through a palladium membrane. These techniques have been evaluated on the basis of initial cost, maintenance, speed and accuracy of analysis, and applicable concentration range. The results of this study showed that the palladium-membrane technique is best suited for routine use. FOR some time investigators have been concerned with the origin, form, and effect of hydrogen in steel. In such stdies', the analysis for hydrogen constitutes one of the most important phases. It is quite apparent that the results for hydrogen concentrations in a given steel are dependent on the method of obtaining the sample, storage of the sample until analysis, preparation of the sample, and analysis of the sample, including all the facets inherent in the calibration and operation of an apparatus for gas analysis. There are a number of means available for determining hydrogen. This is a critical study of some of the more common techniques in use today. In most conventional melting and casting methods, hydrogen concentrations of 4 to 6 parts per million (ppm) in steel are quite common. Because of the undesirable effects of hydrogen on steel there has been increased use of techniques such as vacuum melting,' vacuum casting, and ladle-to-ladle stream degassing, which lower the hydrogen content to levels on the order of 1 to 2 ppm. Therefore, the method used for determining hydrogen in steel must be sensitive and precise. In any analytical procedure for gases in metals there are two distinct operations—the extraction of the gas from the metal and the analysis of the extracted gas. To extract the gas from the steel, three methods have been employed: 1) fusion of the sample with graphite at high temperature; 2) fusion with a flux, such as tin, at a lower temperature; and 3) extraction of the hydrogen from the solid sample at a temperature below the melting point of the steel. Fusion with graphite is the least-acceptable method. The blank in this method is higher and more variable than in either of the other two methods. The hydrogen fraction of the total gas composition usually is between 10 and 50 pct; thus, a larger analytical error is possible. The vacuum-tin fusion4 extraction of hydrogen is probably the most rapid method in use today; the extraction time is usually about 10 min. However, with this system a bake-out of the freshly charged tin for 2 hr is necessary and a change of crucible and a charge of fresh tin are required after each day of operation whether one or thirty samples have been analyzed. In addition, frequent checks of blank rates are required since CO and Na are continually being given up by the steel samples dissolved in the tin bath. The composition of the gas in this method lends itself readily to analysis; although the hydroge concentration may fall to as low as 50 pct, more often it is above 90 pct, thus allowing a more precise analysis (because of less interference from other gases). In 1940 ewell' published the hot-extraction method for extracting hydrogen from the solid sample, comparing analysis for hydrogen extracted at 600°C with similar analysis for the gas extracted at 1700°C by fusion with graphite. Good agreement for hydrogen was obtained between these two methods, provided sufficient time was allowed for extraction at the lower temperature. carsone obtained good results in his comparison of this hot-extraction method with vacuum-tin fusion. Subsequent work by Geller and sun7 and Hill and ohnson' has shown that steel samples should be heated to at least 800°C to effect the release not only of the diffusible hydrogen but also of the "residual" hydrogen that may be present as methane. Since the rate of evolution of hydrogene9l0 depends on such factors as sample size and composition, thermal history, and extent of cold work, a fixed extraction time is not possible. Extraction times of 30 min are normal, but 2 hr are not unusual. Induction or resistance heating may be used in the hot-extraction method. With resistance heating the

Jan 1, 1964

By F. Selig, A. S. Odeh

A second-order approximation to the exact solution of the diffusivity equation corresponding to the pressure build-up of a well producing at a variable rate is derived. This approximation is applicable when the well's shut-in time is larger than the total time elapsed since the well was first produced. The resulting equations are compact in form and easy to use. Thus, the need for Horner's' theoretically precise but rather laborious solution to the above problem is eliminated. In addition, these equations apply where the use of Horner's widely known approximate method is questionable. From a practical point of view, the reported method is best suited for analysis of drill-stem tests and short production tests conducted on new wells. INTRODUCTION The utility of drill-stem and short production tests in reservoir studies has long been recognized by the reservoir engineer. If interpreted correctly they could lead to a wealth of information upon which may depend the success or failure of reservoirs' analyses. Initial reservoir pressure and the average flow capacity are two quantities that are normally sought from a drill-stem and/or a short production test analysis. Pressures are the most valuable and useful data in reservoir engineering. Directly or indirectly, they enter into all phases of reservoir engineering calculations. Therefore, their accurate determination is of utmost importance. The flow capacity kh of the reservoir is indicative of its commercial capability. In addition, it can indicate the presence of a darnaged zone around the wellbore and, thus, the necessity for remedial measures. Of the several methods used to analyze drill-stem and short production tests, Horner's' method is by and large the most common. It applies to an infinite reservoir and or a limited reservoir where the effect of production has not been felt by the boundary. Horner's method makes use of the so-called "point-source" solution of the diffusivity equation. The point-source solution is approximated by a logarithmic function and the superposition theorem is utilized to give the familiar pressure build-up equation where is the shut-in time, q is in reservoir barrels per day and the rest of the symbols conform with AIME nomenclature. Eq. 1 was derived for a well which produced at a constant rate q from time zero to time t and was then shut in. In actuality, such a constant rate of production does not normally obtain. Therefore, a correction must be applied to Eq. 1 to account for the varying rates of production. Horner suggested two methods. The first, which results in a theoretically accurate solution, is rather lengthy and laborious and, thus, it is not suited for routine analysis. The second which has been termed a "good working approximation" is the one used by the majority of the reservoir engineers. In the second method, Eq. 1 is modified by simply introducing a corrected time t, and writing where q is the last established production rate prior to shut-in, and t, is obtained by dividing the total cumulative production by the last established rate. Horner's original paper does not give any indication that this method of correction is based on any theoretical justification. In addition, there is a question as to what constitutes the last established rate. In case of a drill-stem test some engineers use the average rate obtained by dividing the total fluid produced by the total flow time, while others calculate the average rate by dividing the total fluid produced by the last flow-period time. Obviously, different results obtain for the different flow rates used. Because of this, a simple method to the varying-rate case was developed which is theoretically sound and which defines clearly the flow rate and its associated time to be used in the calculations. The final equation arrived at is where q* and t* are a modified rate and time, respectively, and can be easily calculated. In addition, it is shown theoretically that Horner's approximate method, if used for a variable-rate case, gives the correct pressure but would not be expected to give the correct flow capacity. MATHEMATICAL ANALYSIS The general equation governing the flow of slightly compressible fluid in porous media may be written as The elementary solution to Eq. 4, representing an instantaneous withdrawal of Q units volume of fluid at the origin at t = 0, is known as the instantaneous sink

By R. T. Hukki

This paper presents a series of equations for mechanical ball wear, relating parameters of ball size, mill speed, and mill diameter. The fundamental equation, Eq. 12, presented here is introduced to correlate these basic parameters and thus define and clarify the concept of ball wear. This equation is offered as a general rule, which may be modified to apply to individual problems of grinding. BALL wear as observed in grinding installations is the combined result of mechanical wear and corrosion. Corrosion should be a linear function of the ball surface available. Ball corrosion, however, has been studied so little that its effect, although of great importance, cannot be included in the analyses given here. In a separate paper' it is shown that 1 n = 0.7663 np----=— rpm [1] vD P = c, np D kw [2] T=c²(np)n De tph [3] In these equations n — actual mill speed, rpm np = calculated percentage critical speed D = ID of mill in feet P = power required to operate a mill, kw T = capacity of a mill, tph C¹ and c² - appropriate constants in = exponent of numerical value of 1 5 m 1.5 Exponent m is the slope of a straight line on logarithmic paper relating mill speed (on the abscissa) and mill capacity (on the ordinate). It is generally accepted, although not sharply defined, that ball wear in mills running at low (cascading) speeds is a function of the ball surface available. Accordingly, the wear of a single ball may be considered to be a homogeneous, linear function of its surface and of the distance traveled. Thus dw = f¹(d2) . f2(ds) [4] where dw is the wear of a single ball in time dt, d the diameter of the average ball in ball charge, and ds the distance traveled by the ball in time dt. Indicating that ds - a D n dt, the wear of the average ball in time dt becomes dw = f¹(d2) . f2(Dn dt) 1 --- f¹ (d1) f² (D c3 np-----— dt) \/D = c,d² n, D dt The rate of wear of the average ball is given by dw/dt. dw/dt = c, d² np D lb per hr [5] The weight of the ball charge per unit of mill length is a function of D The number of balls of size d in the ball charge is = f³(D2)/f4(d³). The rate of wear of the total ball charge equals the number of balls times rate of wear of the average ball. Thus rate of total ball wear = — . (dw/dt) w. c, . (l/d) . n,, D lb per hr [6] which is the equation of ball wear in low speed mills. In a mill running at a low speed, grinding is the result of rubbing action within the ball mass and between the ball mass and mill liners. When the speed of the mill is gradually increased toward the critical, the impacting effect of freely falling balls becomes increasingly prominent in comparison with the rubbing action. Reduction of ore takes place partly by rubbing, partly by impact. The share of the freely falling balls in the reduction of ore reaches its practical maximum at a speed somewhat less than the critical; at that speed grinding by rubbing has decreased to a low value. It may be reasonable to think that size reduction by freely falling balls should reach its theoretical maximum at the critical speed, if the fall of the balls were not hindered by the shell of the mill beyond the top point; grinding by rubbing would cease at the critical speed. As a first approximation, wear of freely falling balls may be considered to be a homogeneous, linear function of the force at which they strike pieces of rock and other balls at the toe of the ball charge. The force equals mass times acceleration. The mass of a ball is a function of d3 and its acceleration is a function of the peripheral speed of the mill. The wear of a single ball of size d representing the average ball in a ball charge will therefore be w¹ = f3(F) = f (d3) f7 (v). [7] Indicating that v = D n, and n = c³ np 1/vD, Eq. 7 becomes W1 = cn d3 np Do.5 lb per hr. [8] Total wear of the ball charge equals number of balls times the wear of the average ball. Number of

Jan 1, 1955

By M. G. Benz

The superconducting performmce of diffusion-processed Nb3Sn is influenced by its micro structure. High isotropic transverse current density may be achieved in this material by a process which forms a precipitate of ZrO, within the Nb3Sn. FOR an ideal type-I1 superconductor, little or no transport current can be carried in the mixed state; i.e., little or no transport current can be carried above the lower critical field H,,, where the field penetrates abruptly in the form of current vortices or fluxoids, even though full transition to the normal state does not occur until the upper critical field H,,.' Fortunately, nonideal type-I1 superconductors can be readily obtained and these carry large transport currents up to the upper critical field H. Both theoretical and experimental investigations have attributed this current-carrying capability for nonideal type-I1 superconductors to pinning of the fluxoid lattice by heterogeneities in the microstructure of the superconducting material. These heterogeneities may take the form of dislocations or dislocation clusters,2"5 grain boundaries: structural imperfections introduced by phase transformations; radiation damage,8"10 or precipitates.11"15 Nb3Sn formed by diffusion processing is a type-I1 superconductor. Heterogeneities are needed for high superconducting critical currents above H,,. This paper will cover: a) what the microstructure of diffusion-processed NbSn looks like; b) what changes in the microstructure take place when the system is doped with precipitates, and c) how these changes in microstructure influence the superconducting critical currents. EXPERIMENTAL Preparation of Samples. Diffusion processing was used to form the Nb3Sn. The procedure used was as follows: a) coat the surface of a niobium tape with tin; b) heat-treat this tape at a temperature above 930°C to form a layer of Nb3Sn at the Sn-Nb interface. Such a layer of NbsSn is shown in Fig. 1 The thickness of the NbsSn layer formed was controlled by the time and temperature of the heat treatment. The same general procedure was used for preparation of both undoped samples and samples doped with a precipitate. An additional step was included in the preparation of the doped samples which consisted of internal oxidation of zirconium to form ZrOn. The details of the doping process will be reported in a later paper. Sample Testing. The Nb3Sn tape samples were soldered to a copper or brass shunt. Current and voltage leads were then attached to the sample in the usual four-probe resistance measurement configuration. The sample was cooled to 42°K. In some cases it was cooled in the presence of a high magnetic field and in other cases with the field turned off. The results were the same for both cases. The samples were oriented in a configuration with field transverse to current but could be rotated such that the angle between the field vector and the wide side of the tape sample could be changed. Measurements up to 100 kG were done in a superconducting solenoid and measurements above 100 kG in a water-cooled copper magnet at the MIT National Magnet Laboratory. Once the test field was reached, the current in the sample was increased until voltage was detected across the sample. The critical current was taken as the current at which voltage was first detected in excess of background noise. In most cases this was 1 to 2 x 10~6 v for a— in.-wide sample carrying several hundred amperes with a in. separation between voltage leads and with a 10 "-ohm shunt resistance. RESULTS AND DISCUSSION Microstructure. Examination of the microstructure of the undoped Nb3Sn shows rather large-diameter (1 to 2 columnar grains growing outward from the niobium surface toward the tin surface. As the layer is made thicker by longer diffusion times, these grains grow longer. Few new grains are started. Transmission electron microscopy shows little or no second-phase material within the bulk of the Nb3Sn layer. The microstructure of a diffusion-processed NbsSn layer changes quite drastically when the system is doped so as to form a precipitate within the NbsSn layer. Instead of large-diameter columnar grains of NbaSn forming, smaller-diameter (0.5 to 1 ) equiaxed grains of Nb3Sn decorated with the precipitate form. Fig. 2 shows a transmission electron micrograph of a Nb3Sn layer doped with zirconium oxide. This layer has been etched so that one may look between the grains

Jan 1, 1969

By Robert W. Hendricks, John B. Newkirk

A simple method is described for determining the orientation of a single crystal by means of a cylindr cal X-ray camera. Orientation setting to within ±1 deg is attainable by a stereographic analysis of a single cylindrical Laue pattern produced by the originally randomly mounted crystal. Final precision adjustments which permit orientation of the crystal to within ±5 min of arc from the desired position can be made by the method of Weisz and Cole. A chart, originally Presented by Schiebold and schneider7 and which allows a direct reading of the two stereographic polar coordinates of the corresponding pole of a given Laue spot, has been recomputed to aid in the stereographic interpretation of the cylindrical Laue X-ray photograph. Detailed instructions for the use of the chart, a simple example, and a comparison with the conventional flat-film Laue Methods, are presented. 1 HE problem of determining the orientation of the unit cell of a single crystal relative to a set of fixed external reference coordinates is fundamental to most problems of X-ray crystallography and to many experimental studies of the structure-sensitive physical properties of crystalline materials. Several techniques for measuring these orientation relations have been developed which correlate optically observable, orientation-dependent physical properties to the unit cell. Examples of such procedures include the observation of cleavage faces or birefringence, as discussed by bunn,1 or the examination of preferentially formed etch-pits, as discussed by barrett.2 Each of these methods is limited, for various reasons, to an orientation accuracy of approximately ±1 deg—a serious limitation in some experimental studies. Several other limitations decrease the generality of these methods. Of these, perhaps most notable is the absence in many crystals of the physical property necessary for the orientation technique. The most widely used methods for determining crystal orientation are variations of the Laue X-ray diffraction method. Because of the indeterminacy of the X-ray wavelength diffracted to a given spot, the interpretation of Laue photographs is now limited almost exclusively to the procedure of using a chart to determine the angular coordinates of the corresponding pole for each spot. For the flat-film geometries, either a leonhardt3 or a Dunn-Martin4 chart is used in interpreting transmission patterns, whereas a greninger5 chart is used for interpreting back-reflection patterns. A less common method of interpreting flat-film transmission Laue photographs is by comparing the Laue pattern with the Majima and Togino standards,6 or with the revised standards prepared by Dunn and Martin.4 Although this last method is applicable only to crystals with cubic symmetry, it can be very rapid and just as accurate as the graphical methods. The primary limitation of all the X-ray methods mentioned is the relatively small number of Laue spots and zones which are recorded on the flat film. Often, few, if any, major poles appear, thus making interpretation tedious and sometimes uncertain. The use of a cylindrical film eliminates this problem. Schiebold and schneider7 prepared a chart by which the orientation of the specimen crystal could be determined from a cylindrical Laue photograph. However, it was only drawn in 5-deg intervals of each of the two angular variables used to identify the Laue spots, thus limiting the accuracy of orientation to about ±3 deg. An examination of this chart indicated that if it were drawn in 2-deg intervals, crystal orientations to ±1 deg would be attainable. Subsequent use of the replotted chart has confirmed this accuracy. It is the purpose of this paper to describe the redevelopment and use of this chart, and to point out its advantages and limitations. I) CAMERA GEOMETRY AND CHART CALCULATIONS The geometry of the cylindrical camera with a related reference sphere is shown in Fig. 1. The X-ray beam BB&apos; pierces the film at the back-reflection hole B, strikes the crystal at 0, and the transmitted beam leaves the camera at the transmission hole T. One of the diffracted X-rays intersects the film at a Laue spot L. The normal OP to the diffracting plane bisects the angle BOL between the incident and diffracted X-ray beams. The point P on the reference sphere can be located uniquely by the two orthogonal motions 6 and 8 on the two great circles ENWS and BPQT respectively. Because the Bragg angle 8 (= 90 - < BOP) is always less than 90 deg, P always remains in the hemisphere BENWS. Therefore, if every possible pole P is to be recorded on the same stereographic projection, it is necessary that the projection reference point be at T with the projection plane tangent to the sphere at B.* The great

Jan 1, 1963

By A. H. Heim

A method and apparatus for measuring gas porosities of rocks are described. The apparatus can be assembled from commercially available components. In principle, measurements are made by volume substitution at constant pressure. The maximum error is not more than 0.3 porosity per cent. Typical results are given. INTRODUCTION Determining the porosity of rock samples is one of the most important and yet most varied types of measurement in core analysis. Among the many techniques devised are the so-called "gas porosity" methods. An old and well known example is the Washburn-Bunting method.&apos; The U. S. Bureau of Mines2-&apos; described and later improved the apparatus for a now widely used method generally known as the "Boyle&apos;s law" method. In the present form of the Washburn-Bunting method,&apos; the volume of air in the pores of a rock sample at atmospheric pressure is extracted and then collected in a graduated burette at atmospheric pressure. The volume of air is read directly as the pore volume of the sample. The absolute error in reading the collected volume of gas is independent of the total volume; thus, the relative error is larger when the volume is small, as it is for rocks of low porosity. In addition, the sample after measurement contains mercury, which limits its use for other analyses. The Bureau of Mines (or Boyle&apos;s law) method measures directly the solids volume of a sample from which the pore volume and porosity are derived, using a separate measurement of the bulk volume. Gas at a few atmospheres pressure is introduced into a sample chamber of known volume containing the rock sample. The pressure is accurately measured. Following, the gas is expanded into a burette at 1 atm, and the gas volume is read directly. From the initial pressure p, and the final pressure p2 and volume v,, the initial gas volume v1 is calculated using Boyle&apos;s law; that is, p1v1 = p2v2. Volume v, minus the volume of the empty sample chamber is the solids volume of the sample. The accuracy of the method is limited, unless corrections are made, by deviations of the gas from the "ideal" gas-law behavior assumed in the simple form of Boyle&apos;s law. The purpose of the present paper is to describe a method for measuring the gas porosity of a rock which avoids many of these difficulties. Gas volumes are measured directly with the same accuracy as the bulk volumes. Pressures of at least an order of magnitude larger than those of previous methods are employed to insure rapid penetration of the gas into the sample. While special equipment may be built to apply the method, the porosimeter may be constructed as well from commercially available components. For simplicity, the apparatus described will be referred to as the "Constant-Pressure gas porosimeter". THE CONSTANT-PRESSURE METHOD Fig. 1 shows schematically the arrangement of components comprising the present Constant-Pressure porosimeter. Briefly, the method is one of volume substitution and may be considered a null measurement. Omitting (for the present) some of the operational details, the method of measurement consists of the following three steps. 1. After evacuation, the volume of the measuring system (a ballast chamber, a manifold, two gauges and their connections) up to the sample chamber is filled with gas to a high pressure (- 1,000 psi). A sample of the gas at this pressure is trapped in one side of a sensitive differential pressure gauge to serve as the reference pressure for subsequent steps. 2. The evacuated sample chamber containing the rock sample is opened to the measuring system. As the gas expands into the chamber, the resulting decrease in pressure unbalances the differential pressure gauge. 3. The pressure is restored by means of a mercury volumetric pump. The volume of mercury injected exactly equals the free or void volume of the sample chamber (volume of empty chamber minus the solids volume of the rock within). From the injected volume and the known empty chamber volume, the solids volume is obtained and the porosity calculated. The pressure and the volume occupied by the gas are the same before and after opening the sample chamber. Expansion and compression of the gas are incidental operations and do not enter into the calculation of porosity. By the pressure balancing or nulling, the free volume of the sample chamber is merely substituted by an equal and measured volume of mercury. Since the measurements are at constant pressure, there are no compressibility corrections necessary for the sample chamber.

By Robert Baschwitz, John Chipman

The distribution of silicon between liquid silver and Fe-Si-C alloys has been studied at 1420oand 1530°C. The data are consistent with earlier studies. New data of Hager on the liquidus lines of the system Ag-Si and the distribution data are used to obtain the activity coefficient of silicon in both liquid phases. Data on the heat of mixing in iron permit accurate extension to 1600°C. Equilibrium data involving SiO2 and silicon in liquid iron together with revised data on the free energy of SiO2 are used to fix the activity of silicon in the infinitely dilute solution. The binary system exhibits strong negative deviation from ideality. At infinite dilution ? Si at 1600" is 1.25 x 10&apos;3, and at concentrations up to NSi = 0.4 the slope d InySi/dNSi has a constant value of r; = 13. It is found that logysi in the ternary solutzon is approximately but not exactly the same function of Nsi + NC as of NSi in the binary. The results are consistent with currently available data on the free energy of Sic and its solubility in molten iron. LIQUID solutions of the system Fe-Si-C have acquired considerable importance as the laboratory prototypes of blast furnace hot metal. Equilibrium studies involving such solutions and slags approximating those of the blast furnace have yielded useful information concerning the thermodynamic properties of blast furnace slags. In studies of this kind great importance attaches to a knowledge of the thermodynamic activity of silicon in the solution as a function of temperature and composition. An attempt was made by Chipman, Fulton, Gokcen, and askey&apos; to evaluate all of the pertinent data on this system and to deduce the desired relation between activity, composition, and temperature. These authors published data on the solubility of graphite and Sic in molten Fe-C-Si solutions and on the distribution of silicon between liquid iron and liquid silver. They showed further how the activity of silicon in very dilute solutions in liquid iron could be calculated from equilibrium data involving the molten alloy and solid SiO,. These calculations rested on the published thermodynamic properties of SiO, in- cluding its heat of formation which at that time was recorded as -209.8 kcal. This value has been under suspicion for some time and has recently been replaced by the concordant results from two independent laboratories2,3 which place the heat of formation of a-quartz at -217.6 kcal. This revision necessitates a re-evaluation not only of the activity of SiO2 in slag but also of silicon in molten iron. It is the purpose of this paper, therefore, to recalculate the activity of silicon, and in furtherance of this objective to present new data on its distribution between liquid Fe-Si-C alloys and liquid silver. HEAT OF SOLUTION OF SILICON IN IRON In order to determine the effect of temperature upon the activity coefficient it is necessary to know the heat of solution of silicon in iron as a function of composition. This is found in the data of Korber and Oelsen4 shown in Fig. 1. The curve corresponds to the following equation, which is of a form suggested by Wagner:5 Here AH is the heat absorbed in kilocalories in forming one gram atom of molten alloy from its molten elements and the N&apos;s are atom fractions. The relative partial molal enthalpies of the components, each referred to its pureliquid state and defined as zFe = aFe - PFe and zsi = HSi — -psi, are shown graphically. At low concentrations zSi = -28.5 kcal, in agreement with Kijrber and Oelsen&apos;s computation. This is in good agreement with the value of -29.3 kcal obtained by Chipman and Grant6 using an entirely different method. ACTIVITY AT INFINITE DILUTION From the known free energy of SiO, it is possible to obtain the activity of silicon in dilute solution in liquid iron from equilibrium studies. The heat of formation of a-quartz is —217.6 kcal and the heat capacity and entropy data are given by Kelley and ~ing.&apos; The free energy of formation of ß-cristo-balite at temperatures above the melting point of silicon is expressed by the following equation: Si(Z) + O2(g) = SiO2 (crist); AF" =-226,500 + 47.50T [I] The value of the deoxidation product for silicon [%Si] x [%O]2 at 1600°C according to Gokcen and chipmans is 2.8 x 10"5, in agreement with results of Hilty and Crafts.9 More recent works of Matoba,

Jan 1, 1963

By P. C. J. Gallagher

A number of recmt determinations for the ratio of extrinsic to intrinsic stacking fault energy in fcc solid solutions are examined. Some of these arise from incomplete analyses which can yield only approxi?nate values for the ratio. Reliable results, on the other hand, obtained using extrinsic-intrinsic fault pairs, show that the extrinsic and intrinsic fault energies are essentially equal in several materials. There is some reason to believe that this finding is of general applicability to fcc elements and alloys. A wide range of values has been obtained for the relative magnitudes of the extrinsic and intrinsic stacking fault energies (yext and yint, respectively) in recently published studies in a variety of materials. In contrast, Hirth and Lothe&apos; using a central force model have shown that out to tenth nearest-neighbor interactions the perturbation in energy caused by both types of fault is the same. Although the model used is not completely valid in metals, there is nevertheless some indication that marked variations of yext/nnt should not be observed from one material to another. In early work in Cu-A1, Cu-Ge, Ni-Co, and stainless steel all the deformation faults observed in the electron microscope were found to be intrinsic in nature, which led to an attitude that the extrinsic fault energy must be considerably greater than the intrinsic. Extrinsic faulting arising from deformation has, however, more recently been directly observed in Au-4.8 at. pct n;~ Ag-6 at. pct Sn and Ag-8 at. pct sn; Ag-7.5 at. pct In and Ag-11.8 at. pct 1n;"&apos; pure silwer and Ag-0.5 at. pct ~n;&apos; and Cu-22 at. pct Zn, Cu-30 at. pct Zn, and Cu-7.5 at. pct ~1.l&apos; Multilayer loops containing extrinsic faulting have also been observed in quenched aluminum." While peak asymmetries in X-ray faulting probability studies were generally attributed to the presence of twins,Lelel2 has recently reinterpreted earlier X-ray data in Ag-Sb alloysU in terms of the presence of extrinsic faulting. The determinations of yeXt/yint made from the above studies are shown in Table I, with a brief description of the techniques employed. A number of the methods utilized are deficient in one or more respects, and the reliability of the values listed will be discussed. ~ele&apos;~ recognizes that his approximate determinations of yext/yint assumes equal numbers of extrin-sically and intrinsically faulted dislocations. It is well-known, however, that such an assumption is not at all justified since extrinsic faulting has but rarely been observed in samples studied in the electron microscope. The only conclusion that should be drawn from the X-ray results at present is that the total intrinsic scattering cross section (i.e., the product of the width of the intrinsically faulted dislocations with their density) is approximately ten times greater than for extrinsic faulting in these particular samples. An important point is that the relative magnitudes of the energies cannot be inferred from results of this type, unless the intrinsic and extrinsic faults form with equal ease. One must recognize that, although a formation barrier may restrict the amount of extrinsic faulting which occurs, this in no way implies that the extrinsic and intrinsic energies should be different. It is unlikely that a worthwhile estimate of the relative densities of extrinsically and intrinsically faulted dislocations can be made at the high deformations present in X-ray samples. ~oretto,&apos;~ from a statistical argument applied to the nonobservation of extrinsically faulted tetrahedra out of a large sample, concluded that yeXt/yint could not be less than -4.5. However, the present author feels that a high-energy formation barrier as just supposed also explains this finding satisfactorily and that no conclusion can possibly be drawn concerning the actual extrinsic stacking fault energy. The same argument also serves to explain the fact that extrinsic faulting has been relatively little observed in the electron microscope. Extrinsic-intrinsic node pairs and isolated extrinsic nodes were first reported by Loretto~ and subsequently by Ives and Ruff,&apos; Gallaher,&apos; and Gallagher and Wash-burn.&apos; Ives and Ruff&apos; found a wide spread in the ratio of extrinsically to intrinsically faulted area in the node pairs they observed, and drew the very tentative conclusion that yeXt/yint 2 2. They recognized that a straightforward comparison of the size of the faulted areas could provide no more than a qualitative result without a theoretical analysis of the dislocation geometry associated with extrinsic faulting. A theoretical

Jan 1, 1969

By L. C. Chang, T. A. Read

Diffusionless transformation in Au-Cd single crystals containing about 50 atomic pet Cd was investigated by means of X-ray analysis of the orientation relationships, electrical resistivity measurements, and motion picture studies of the movement of boundaries between the new and old phases during transformation. The nucleation of diffusionless transformation by imperfections and the generation of imperfections by diffusionless transformation were discussed. THAT connections exist between plastic deformation and diffusionless phase changes has long been recognized. Thus it is often possible to produce a diffusionless phase change in a temperature range, above that in which the change occurs spontaneously, by cold-working the initial phase. Certain aspects of the dislocation theory of the plastic deformation of crystalline solids also provide for a rather direct connection between the processes involved in plastic deformation and in diffusionless phase changes. Heidenreich and Shockleyl have pointed out that simple edge dislocations in f.c.c. metals are probably unstable, and that the more probable lattice imperfections, called extended edge dislocations, consist of two half dislocations separated by a distance of the order of magnitude of 100A. The region about two atomic planes thick between the half dislocations because of this stacking fault may be described as having the hexagonal close-packed structure. Presumably the stacking faults observed by Barrett" fter cold-working f.c.c. Cu-Si alloys resulted from the passage of such half dislocations through the lattice of the initial phase. It is now becoming clear that the development of a detailed theory of the atomic movements involved in diffusionless phase changes will require a consideration of the role played by lattice imperfections, just as such considerations are necessary to the understanding of plastic deformation mechanisms. This point of view has been recently set forth, for example, by Cohen, Machlin, and Paranjpe3 who pointed out the role which might be played by screw dislocations in nucleating diffusionless phase changes. The present paper reports on some aspects of the diffusionless phase change in single crystals of the beta phase alloy Au-Cd which serve to emphasize further the importance of lattice imperfections in diffusionless phase changes. The diffusionless phase change of Au-Cd possesses several remarkable features. One of these is that the interface between the high-temperature beta phase and the low-temperature orthorhombic phase typically moves with a low velocity, in contrast to the behavior observed in the transformation of austenite to martensite. Motion pictures of this slow interface motion have been prepared in the course of the work reported here. Another important feature of the Au-Cd transformation is the small amount of undercooling observed. The reverse transformation occurs on reheating to a temperature only 20" higher than the transformation temperature observed on cooling, and under some circumstances the hysteresis observed is substantially less than this. This narrow temperature range between transformation on heating and cooling is presumably in part a consequence of the fact that the transformation requires a lattice shear of only about 3". Finally, the orthorhombic product phase possesses unusual mechanical properties, as was first pointed out by olander&apos; and Benedicks." After completion of the transformation on cooling the specimen can be severely deformed, yet on the release of load it springs back to its original shape in a rubber-like manner. Explanation of this phenomenon will require an understanding of the lattice imperfections in the orthorhombic structure and, correspondingly, of those in the initial body-centered cubic structure. Single crystals of Au-Cd alloy containing 47.5 and 49.0 atomic pct Cd were prepared from fine gold (99.95 pct purity) and chemically pure cadmium (99.99 pct purity) by melting the alloy in an evacuated and sealed fused quartz tubing and growing into single-crystal form by the Bridgman method. The Au-Cd alloy containing 47.5 atomic pct Cd undergoes a diffusionless transformation from an ordered body-centered cubic structure to an orthorhombic structure when it is cooled to about 60°C, while the reverse transformation takes place when the alloy is heated to about 80°C, according to electrical resistivity studies. The structures of these two phases have been studied by Blander,4 reinvestigated by Bystrom and Almin.e he lines of Debye photo-gram of powdered samples of Au-Cd alloy containing 47.5 atomic pct Cd prepared in this laboratory were identified and agreed fairly well with those of

Jan 1, 1952

By M. D. Arnold, H. J. Gonzalez, P. B. Crawford

A method is presented for estimating the effective directional permeability ratio and the direction of maximum and minimum permeabilities in anisotropic oil reservoirs. The method is based on the principle that production from a well in an anisotropic reservoir results in elliptical isopo-tentials about the well, rather than circular. Bottom-hole pressure data from three observation wells surrounding a producing well are required to apply the method. The method involves fitting field pressure data to a set of general charts of isopotentials and making a few simple calculations until a solution is found. The method is based on a steady-state equation for homogeneorrs fluid pow. In addition to the method, a brief discussion of the theory underlying it is presented. INTRODUCTION The existence of a different permeability in one direction than another in oil reservoirs has been mentioned in several papers. Hutchinson&apos; reported laboratory tests on 10 limestone cores and pointed out that one-half of them showed significant, preferential, directional permeability ratios, the average being about 16:1. Johnson and Hughesz reported a permeability trend in the Bradford field in the northeast-southwest direction with flow being 25 to 30 per cent greater in that direction. Barfield, Jordan and Moore -eported an effective permeability ratio of 144:1 in the Spraberry. Crawford and Landrum4 showed that sweep efficiencies could often vary by a factor of two to four, and sometimes considerably more, due to variations in flooding direction and patterns in anisotropic media. These findings indicate that the poss&apos;bility of anisotropy may be worthy of consideration in the development of an oil field. In considering this, it should first be determined if anisotropy exists. If it does, the direction of the maximum and minimum permeabilities and the ratio of their magnitudes are quantities which can be of value in planning the most efficient well-spacing patterns. Past methods of determining these quantities have included analysis of oriented cores and analysis of flooding performance of pilot injection patterns. In recent work, Elkins and Skov5 resented an analysis of the pressure behavior in the Spraberry which accounted for anisotropic permeability. This work was based on the transient pres- sure distribution in a porous and permeable medium, with the solution expressed as an exponential integral function involving rock and fluid properties. The purpose of this study is to provide a method, based on steady-state equations, of estimating the direction and relative magnitude of permeabilities in an oil reservoir from field pressure data and well locations only. The method presented is based on work by Muskat6 which shows that Laplace&apos;s equation represents the steady-state pressure distribution for homogeneous fluid flow in homogeneous, anisotropic media if the co-ordinates of the system are shrunk or expanded by replacing x with it is desirable that data be obtained early in the history of a field because knowledge of an anisotropic condition would allow new wells to be spaced in such a manner that reservoir development and subsequent secondary recovery programs could be planned more efficiently. THEORETICAL CONSIDERATIONS A brief discussion of the theoretical basis on which the graphical solution was developed is presented in this section. Muskat&apos;s two-dimensional6 olution for the pressure distribution in an homogeneous, anisotropic medium with an homogeneous fluid flowing can be algebraically manipulated to show that the isobaric lines are perfect ellipses. The ratio of the major axis to the minor axis, a/b, is related to the permeability ratio, k,/k,, as follows. alb = dk,/k,--...........(1) It can also be shown that the pressure varies linearly with the logarithm of the radial distance from the producing well. However, the gradient along any ray is a function of the orientation of that ray, and a ..xiable is present when anisotropy exists which cancels out for a radial (isotropic) system. For a system such as that described, a dimensionless pressure-drop ratio was developed which is completely independent of the actual magnitude of the pressures. This was done by arranging Muskat&apos;s solution in such a way that aIl variables cancelled out except k,/k, and well positions. However, this solution depends on having a co-ordinate system with axes coinciding with the major and minor axes of the elliptical isobars. Thus, it was necessary to introduce a co-ordinate system rotation factor. The two unknown variables are then k,/k. and 0, and the two measured dimensionless pressure-drop ratios are related to the unknown variables as follows.

By M. I. Sherman, J. D. H. Strickland

USE of a hydrometallurgical approach to the oxidation of sulfide ores and extraction of metals therefrom may have advantages over the more common smelting techniques when a low grade deposit is difficult to concentrate or the subsequent separation of metals, coexisting in the ore, is laborious by any known smelting operation. For economic reasons, the most promising oxidants are either atmospheric oxygen or electric power. The use of oxygen, or air under pressure, has recently been revised. Pyrrhotite has been converted to iron oxide and elementary sulfur&apos; and a variety of sulfides have been treated by Forward and co-workers.2-4 Generally sulfate is the end form of the sulfur but with galena in an acid medium, elementary sulfur can be formed." For economic reasons chlorine and ferric iron salts are about the only possible alternatives to the atmosphere as oxidizing agents for base metal sulfides. If aqueous solutions of chlorine or ferric iron are employed, the reduction products can be oxidized electrolytically in situ and used again, thus acting as catalysts for electric power as oxidant. The use of ferric salts for this purpose is established hydrometallurgical practicea but, although chlorine gas has been employed in the dry state at an elevated temperature, its use in aqueous solution at or near room temperature has not found favor. The reaction of chlorine water with the soluble sulfide ion has been studied by several workers,7-9 and both sulfate and elemental sulfur are found as end products, the latter being favored by the presence of a low concentration of oxidant relative to that of sulfide in solutions of about pH 9 to 10. Of direct bearing on the work in hand are an early American patent" and a recent Austrian patent." The former advocates stirring powdered ore with an aqueous solution of ferric chloride chlorine oxides and chlorine. In the latter it is claimed that both metal and sulfur can be obtained by electrolysis, in a diaphragm cell, of a metal ore slurry in brine. Details in these patents are scant and no data or explanation is given for the mechanism of the reaction which, in the Austrian work, is attributed to the (unlikely) action of nascent chlorine at the anode surface. No mention is made of possible differences in behaviour between various ores. Apparatus A complication encountered when working with chlorine water is that a serious loss of chlorine occurs by gas partitioning unless an enclosed system is used and any air space in the apparatus is kept very small and constant. Arrangements were made, therefore, to take out samples for analysis without letting air into the system to replace the liquid removed. For convenience in studying a heterogeneous reaction the apparatus was so designed that a reproducible controlled stirring rate could be maintained and the ratio of surface area of ore to volume of solution was approximately constant throughout any experiment. The apparatus used is shown in Fig. 1. The ground ore was placed in the horizontal cylindrical vessel, A, of about 1 liter capacity, heated by a constant temperature circulating bath pumping water through the concentric jacket, B. By adding chro-mate to this water, an ultraviolet radiation filter effectively surrounded the reaction vessel, greatly reducing any possible photochemical decomposition of chlorine solutions. Stirring was effected by glass paddles, C, attached by an axle to a magnet which was rotated by another powerful Alnico magnet, D, outside the glass end, this magnet being itself rotated by an electric motor electronically controlled to constant speed. Speed could be varied from about 150 to 900 rpm and was measured and held to within 1 pct of a given value. The end of the reaction vessel remote from the stirring magnet was closed by another one-ended glass cylinder, E, connected by thin polyethylene bellows, F, clamped by screw clamps and watertight rubber gaskets to the main vessel. Through E, a glass electrode and calomel electrode projected into the solution and a hypodermic syringe pierced a small bung and allowed acid or alkaline to be added to maintain a constant pH. By pushing the fully extended bellows until the two cylinders touched, from 50 to 100 ml of solution could be forced out through a sintered disk into the three-way tap system, G, either to waste (for flushing purposes) or up into a 10 ml burette where the solution could subsequently be measured out for analysis. The ore samples were introduced at H, the tube being stoppered by a thermometer of —1 to +52ºC range, graduated to 0.1°C intervals. To prevent ore from being ground in the end bearings of the stirrer these bearings were pro-

Jan 1, 1958

By D. E. Spiller

Introduction Precious metals recovery from ore can generally be accomplished using gravity concentration, flotation, and/or hydrometallurgical (leaching) techniques. The objective of this paper is to show why gravity concentration can be an important part of recovery systems that employ leaching as the primary unit operation. A brief discussion of modem gravity concentration equipment is also presented. Discussion Gravity concentration of ores has generated increasing interest in recent years. Reasons for this interest include: • Gravity concentration is environmentally attractive. There is little or no use of reagents. Hence, it is relatively nonpolluting. • The cost of cyanide has continued to increase. Therefore, cost savings may be realized whenever leaching feed tonnage can be reduced by preconcentration. • Compared to flotation and leaching, gravity equipment costs are low per processed ton. Field installation costs for gravity circuits usually are less because many&apos; units are supplied as self-contained modules. Also, the cost required to supply services, particularly power, to a gravity plant site are also less. In situations where preconcentration at coarse particle size is applicable, significant grinding equipment savings may be possible. • Gravity circuit operating costs are also relatively low compared to typical flotation and leaching circuits. Reagents, power, maintenance, and manpower savings in a well-engineered gravity plant may be realized. Again, if grinding is reduced, significant power and steel (media and liners) savings are possible. •In recent years, more efficient gravity concentrating devices have been developed. Benefits to Precious Metal Leaching Gravity beneficiation can complement precious metal leaching in two ways. First, the recovery of coarse liberated values before leaching may reduce leach time requirements and may reduce reagent consumption. Second, gravity preconcentration can reduce the size of a leach plant by decreasing the quantity of material to be leached. Coarse gold and silver have been shown to leach rather slowly. Kameda (1949) and Habashi (1967) have investigated the kinetics of cyanide leaching systems. They concur that in a heterogeneous reaction, the rate of gold and silver dissolution is directly proportional to the surface area. Thus, the instantaneous rate of dissolution for spherical 0.37 mm (400 mesh) gold is theoretically -25 times faster than for the same amount of gold at .841 mm (20 mesh), based on data from Fuerstenau, Chander, and Abouzeid (1979). Conversely, coarse liberated, +.841 mm (20 mesh), gold is more readily recovered by gravity concentration than is fine, -.037 mm (400 mesh) gold. Therefore, it is apparent that the two recovery systems complement one another. Figure 1 data demonstrates the potential synergism. A sample of - 3.327 mm (6 mesh) Nevada gold-bearing ore was cyanide leached using conventional bottle-roll test procedures. Gold extraction was determined as a function of leaching time. A second sample split from the same leach feed material was hand jigged to remove a coarse heavy mineral fraction, including virtually all of the +.210 mm (65 mesh) liberated free gold. This second sample, with the coarse gold and heavy minerals removed, was subjected to an identical cyanide leach procedure. Figure 1 presents the resulting comparative extraction data. Note that the percent gold extraction for the sample containing no +.210 mm (65 mesh) free gold includes the coarse gold recovered by gravity. The data show that the sample containing coarse gold required about 72 hours of leaching time to achieve 80% extraction. This compared to about 22 hours of leaching time for 80% gold recovery from the sample that contained only -.210 mm (65 mesh) free gold. Thus, there was a 69% reduction in leaching time. The improved extraction data is not wholly attributable to coarse gold removal, but rather it was the combination of gold removal and rejection of other heavy mineral cyanide consumers or leach retardants. Further investigation was not warranted at this time. Preconcentration is the second manner in which leaching systems can benefit from gravity concentration. The premise is that preconcentration can reduce the quantity of leach feed, which, in turn, may reduce leaching costs. Figure 2 presents preliminary data developed by CSMRI for US Minerals Exploration (USMX). Centennial Exploration Inc., in agreement with USMX, is proceeding with evaluations to determine the suitability of various processing schemes for recovery of gold values from the Montana Tunnels property. The data shows how the ore can be preconcentrated by gravity techniques to result in a reduced feed tonnage to secondary extraction techniques, presumably flotation or cyanide leaching. Testing has shown that Reichert cones, followed by treating the cone concentrate on spirals, can deliver about 88% gold recovery in about 13% weight, that is, 87% weight rejection. Consequently, fine grinding and reagent costs are attributable to only 13% of the plant feed rate. Cost data is not yet available, but the potential exists for significant cost savings.

Jan 1, 1985

By R. A. Wood, R. I. Jaffee, H. R. Ogden, D. N. Williams

The tensile properties, creep resistance. and thermal stability of highly alloyed Ti-Al-Zr alloys were examined. On the basis of these studies, the Ti-7Al-1ZZr composition was selected for more complete evaluation. The alloy was found to be weldable and free from excessive directionality. In addition, it developed maximum properties without requiring heat treatment other than an annealing operation in the alpha field. The alloy was recommended for scale up and is presently being investigated on a production-level basis. One of the more attractive properties of titanium alloys is their ability to withstand stress at moderately high temperatures, and a considerable amount of effort has been devoted to increasing the maximum service temperature of titanium alloys. This work has suggested that the optimum alloys for high-temperature service will be single-phase a (close-packed hexagonal) alloys containing significant amounts of aluminum. However, the maximum amount of aluminum which can be alloyed with titanium is between 6 and 8 pct,l since at high-aluminum contents an embrittlement reaction occurs in the anticipated service temperature range, 800" to 1100°F. It has been shown that the embrittlement reaction involves decomposition of the high-aluminum a phase to one or more new phases.&apos; Since this reaction does not occur at intermediate or low-aluminum contents, it was felt that intermediate Ti-A1 alloys might be strengthened by a-soluble ternary additions without inducing the embrittlement reaction. The first alloying addition considered was tin, which shows extensive solubility in a titanium and has moderate strengthening tendencies. Unfortunately, it was soon apparent that tin also promoted the embrittlement reaction, and that to obtain a stable alloy, the aluminum content had to be reduced as the tin content was increased. The second alloying addition considered was zirconium, which is similar to tin in its effects on titanium. This element did not contribute to the embrittlement reaction and, in fact, appeared to increase the maximum amount of aluminum which could be alloyed with titanium without inducing instability. This paper describes an investigation of the Ti-A1-Zr a alloy region. Alloys containing from 4 to 12 pct A1 and from 6 to 15 pct Zr were examined. The properties of these alloys are described and the bases for selecting an optimum composition is outlined. This composition, Ti-7A1-12Zr, is presently being scaled up in tonnage quantities, and is being evaluated extensively throughout the industry. In addition to presenting the basis for its selection, this paper presents a description of the properties developed in laboratory material as determined during the alloy investigation. These properties suggest that this alloy can fill an important position in applications requiring light weight, fabrica-bility, weldability, and strength to 1000oF or higher. EXPERIMENTAL PROCEDURES Titanium alloy ingots were prepared by inert electrode arc melting under an argon atmosphere. Alloying elements used were 110 Bhn titanium sponge, high-purity aluminum, and reactor-grade zirconium. Pancake-shaped ingots were prepared weighing approximately 300 g. The composition of the ingots was checked by weight measurements before and after melting. The pancake ingots were forged at 2000°F to approximately half their original thickness to give a flat plate roughly 1/2 in. thick. This plate was then rolled at 1800&apos; to 1600°F to 0.250 in. thick. All of the alloys examined fabricated well. However, alloys containing 15 pct Zr tended to overheat due to exothermic oxidation, and scaling was excessive. As might be anticipated from its effect in decreasing the ß transus, increased zirconium appeared to improve fabricability somewhat, especially during rolling at lower temperatures. Except for a limited study of heat-treatment response, all alloys were examined in the a-annealed condition. Prior to heat treatment the a and ß tran-sus temperatures were determined by metallo-graphic examination of samples quenched after annealing at 50-deg intervals in the transformation region. These data are shown in Fig. 1. Recrystal-lization appeared to occur in about 1 hr in the range 1300º to 1500ºF. Therefore, alloys were annealed for 1 hr at 1550ºF (4 and 5 pct Al), 1600ºF (6 through 7-1/2 pct Al), or 1650°F (8 or more pct Al). This produced an equiaxed a grain structure. In most alloys, a "ghost" structure was visible after the a-annealing treatment, as shown in Fig. 2. This structure apparently resulted from the acicular

Jan 1, 1963

By R. I. Jaffee, R. M. Evans

IN recent years, the interest in liquid metals as heat-transfer media for power plants has been very great. The possibility of the development of nuclear power plants has increased this interest and served as the impetus behind much research on low melting metals and alloys for such purposes. The principal reasons for consideration of liquid metals as heat-transfer media lie in their high thermal conductivity and consequent high heat-transfer coefficients, stability at high temperatures, and the high ranges of temperature possible. The element gallium possesses some of the requisite properties for a heat-transfer liquid. It is a unique material, having a low melting point and a high boiling point. Pure gallium melts at 29.78oC, and suitable alloying will produce a metal which melts below room temperature. The boiling point is about 2000°C. As it is a liquid metal, the heat-transfer characteristics would be good. Gallium is not now readily available, due in part to a lack of uses for the metal. Nevertheless, it is not a rare element, and a sufficient supply of gallium exists to permit its consideration for this use. Since gallium has some promise as a heat-transfer liquid, owing to its unique properties, research on the subject was carried on at Battelle Memorial Institute at the request of the Bureau of Ships, U.S.N. The research had as its objectives the determination of the effect of alloying on the melting point of gallium, and the study of the corrosion of possible container materials. In this research, alloys were found which had significantly lower melting points than pure gallium, but none which simultaneously fulfilled other additional requirements, chiefly the corrosion problem. Neither was it found possible to reduce the melting point of certain otherwise suitable alloys appreciably by small additions of gallium or gallium alloys. The results gave little hope that gallium alloys can be developed which enhance the good properties and minimize the undesirable characteristics of elemental gallium. Thus, gallium now appears less promising than other metallic heat-transfer media. The experimental thermal-analysis techniques used in this work have been described.&apos; Experimental Results As a first approximation, the development of low melting gallium alloys was based on alloying elements suitable for use in a nuclear power plant, which also lowered the melting point of gallium. Information from the literature, summarized in Table I, indicates that. tin, aluminum, and zinc are the only suitable elements which cause a lowering of the melting point of gallium. Indium and silver also lower the melting point of gallium, but are of little interest for use in nuclear power plants. Of the elements reported not to lower the melting point of gallium, there is some ambiguity on the behavior of copper. Weibke3 obtained solidus arrest temperatures of 29°C for Cu-Ga alloys from 60 to 90 pct Ga, 0.8C lower than the generally accepted melting point. This may be the effect of a eutectic close to gallium, or, more likely, the result of impurities, or experimental error. The seven elements listed in Table I whose effects were not known were of potential interest if they lowered the melting point of gallium. Their effects were determined experimentally for this reason. Binary alloys containing nominally 2 pct of each of these elements were prepared in the form of 2-g melts by placing the components in a graphite crucible and holding them in an argon atmosphere at 370°C for 5 hr. These melts were then subjected to thermal analysis. In all cases. the solidus temperature was the melting point of gallium. Since these elements (As, Ca, Ce, Mg. Sb, Si, and T1) did not lower the melting point of gallium, they were not considered further as components of a eutectic-type alloy. Ga-Sn-Zn Alloys Preliminary considerations of this system for low-melting alloys were encouraging. All three binary systems were of the simple eutectic type. The composition and melting points of the eutectics were as follows: Sn-9 pct Zn (199°C), Ga-8 pct Sn (20°C), and Ga-5 pct Zn (25°C). Therefore, the probability of a ternary eutectic was high. For reasons to be discussed later, aluminum could not be used as an alloying constituent, leaving the Ga-Sn-Zn system as the only one of interest for low-melting gallium-

Jan 1, 1953

By G. H. Wilson

INTRODUCED in 1946 to serve a need in power-plant operation, closed-circuit TV has been used by well over 200 organizations in approximately 25 different industries. Known as industrial television, or simply ITV, it can be described as a private system wherein the television signal is restricted in distribution, usually by confinement within coaxial cable that directly connects the TV camera to one or several monitors, Figs. 1, 2. The picture is continuous and transmission is instantaneous, permitting an observer to see an operation that may be too distant, too inaccessible, or too dangerous to be viewed directly. Destructive testing or the machining of high explosives can now be conducted hundreds of feet away by personnel who still have close control through the eyes of the TV camera. It is also possible for one man to control operations formerly requiring the co-ordinated efforts of several workers. For example, at a large midwestern cement plant conveyance of limestone from primary crusher to raw mill and loading into five storage bins once necessitated the work of two men, one having little to do but prevent spilling of material by manually moving the tripper on the belt conveyor as occasion required. TV cameras mounted on the tripper now provide bin level indication to the conveyor operator at the crusher position so he is able to control the entire loading operation remotely, Fig. 3. By means of a switch, the picture from either camera is alternately available on a single viewer, or monitor, Fig. 4. Each camera is mounted on the tripper by means of a simple adjustable support and looks down into the bin, which is identified by the number of cross members on the vertical rod. Each associated power unit is located on a platform above the camera, Fig. 5. This centralized control by means of TV often has produced superior results, and in many instances saving in operating costs has been sufficient to write off equipment costs within six months to a year. Where a key portion of a process may be enclosed or otherwise inaccessible, TV again reduces the likelihood of mistakes and permits closer control by making available to the operator valuable information he might otherwise never possess. An example of this can be found at a strip mine where the coal seam lies 50 ft or more below the overburden, which is removed by a large wheel shovel. From his centrally located position the shove1 operator was unable to judge accurately to what extent the wheel buckets engaged the earth. His chief indication of efficiency was the amount of overburden on the belt conveyor as it passed his control point 75 ft from the wheel. Now, two television cameras mounted on the tip of the boom permit the operator to view the wheel from each side and provide him with a close-up view of the buckets so that he can take immediate and continuous advantage of their capacity, quickly compensating for ground irregularities and avoiding obstructions, Fig. 6. While the word television conjures up visions of highly complex and intricate apparatus such as that employed in modern TV studios and transmitting stations, the term industrial television should indicate compact, straightforward equipment. Most present-day ITV systems contain fewer than 25 tubes including camera and picture tubes. The average home television receiver alone requires at least that many tubes. Equipment like that illustrated in Fig. 1 contains only 17 tubes, of which 3 are in the camera. It can operate continuously and dependably, without protection, in any temperature from 0" to 150°F. It consumes less current than a toaster and weighs under 140 lb. Camera and monitor may be separated by 1500 to 2000 ft and by greater distance with additional amplification. This equipment is designed to withstand vibrations up to 21/16 in. and will operate successfully under more severe conditions of vibration and heat when suitable enclosures are provided. Any number of cameras may be switched to a single monitor, and any number of monitors, within reason, used simultaneously. Two types of applications in the mining industry have already been described. A third under serious consideration by several organizations will make use of ITV for remote observation of conveyor transfer points at copper concentrating plants so that evidence of belt breakdown and plugging of transfer chutes can be spotted immediately and costly overflow of material avoided. A television camera will soon be installed to view a trough conveyor near the exit of an iron-ore crusher to indicate clogging of the crusher as evidenced by reduction or absence of material on the

Jan 1, 1955