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By A. E. Austin, J. R. Doig

Ti-Cr alloys age harden after B-quenching by formation of an W-phase. The structure and orientation of this transition phase have been determined. The hardness ap-pears to be caused by strain in the B-matrix produced by concurrent enrichment of alloy content in the W-phase. EARLY studies of age hardening in B-stabilized titanium alloys revealed the formation of a transition phase named o' This phase was detected by X-ray diffraction in binary alloys of Ti-Fe, Ti-Cr,

Jan 1, 1958

By D. J. Carney, J. Chipman, N. J. Grant

An absolute calibration has been achieved for sampling and analyzing liquid steel for hydrogen based on Sieverts' values of hydrogen solubility in iron. Further checks were made in nickel, iron-nickel, and 18-8 stainless melts. The sampling method was successfully applied to a large number of commercial steel melts in various types of furnaces. The concurrent problem of sample storage was also solved. THE problem of sampling of liquid steel for hydrogen is more difficult than the problem of analysis. Hydrogen has a greater mobility than any other element; it is the only element that will diffuse out of steel at room temperature. In accordance with the known relation between temperature and diffusivity, the diffusion of hydrogen is extremely rapid at the temperature of liquid steel. Consequently, in the past, attempts to quench molten steel to retain the dissolved hydrogen have not been as successful as similar attempts with nitrogen or oxygen. The retention of hydrogen is especially difficult since it will continue to escape at room temperature even if the molten metal has been rapidly quenched. The hydrogen sampling methods may be classified broadly as (1) the rapid quenching techniques which attempt to preserve super saturation and (2) those methods which attempt to collect the gases evolved as the metal freezes and cools to room temperature. With all methods of sampling but especially with the rapid quenching methods, there is also a concurrent problem of storing the sample until analysis is undertaken. The two problems are interrelated and must be considered together when discussing liquid steel sampling techniques. The major obstacle in liquid steel sampling has been that there was no way of evaluating properly the various sampling and storage techniques which have been developed. It has not been possible to compare the accuracy of the various methods. For evaluation of sampling methods some known standard is required; this may be found in the published work on the equilibrium solubility of hydrogen in liquid iron which has been determined in three separate investigations.1,2,3 The data are shown in table I. Sieverts' original method of obtaining gas solubilities involved measurement of the volume of gas required to saturate the liquid metal contained in an evacuated bulb of known volume. This procedure, improved by introduction of high frequency induction heating, has yielded reproducible results which appear to be entirely dependable. Using Sieverts' method, it has been proved for iron and other metals and alloys that the hydrogen solubility is proportional to the square root of the pressure of the hydrogen gas above the liquid, so that one may work with various partial pressures of hydrogen to put different amounts of hydrogen in solution. From Sieverts' law it is possible to know accurately what amount of hydrogen is dissolved in the liquid metal at the time of sampling. The accuracy of a sampling method may be gauged by reference to this known solubility. It was the purpose of this investigation to develop a method for sampling liquid steel and for storing the sample so that the analytical result would represent the actual hydrogen content of the metal bath. For this purpose an induction furnace was

Jan 1, 1951

By E. Furubayashi, T. Taoka, S. Takeuchi

Recrystallization textures developed from artificial nucleation around indentations were compared with spontaneously recrystallized textures in two rolled 3 pet Si-Fe single crystals with (111)[112] and (100) [011] orientations. Orientation of individual grains in some stages of their growth was also directly determined by a microetch-pit method. In both the crystals, the recrystallization texture was characteristic of each orientation regardless of any nucleation conditions, artificial or spontaneous, though artificially nucleated grains were more randomly oriented in finer stage. These results support the concept that the process of selection of growing grains controls the formation of the prejerred recrystallization texture. Discussion is given on the factors controlling the mobility of grain boundaries. Recrystallization textures are formed through a number of stages which are customarily called by the terms recovery, nucleation, and grain growth. Many experiments have been done to clarify the mech- anism of formation of recrystallization texture but no fully successful explanation has been proposed. The texture consists of oriented grains which nucleate at many sites of the deformed matrix and grow at the expense of the matrix during annealing. The current theories on the formation of recrystallization texture are based on the following two extreme hypotheses: 1) the oriented nucleation theory by Burgers et al.;1, 2 2) the oriented growth theory by Beck et al.3-5

Jan 1, 1968

By F. M. Lewis, J. F. Myers

W. I. Garms-—The authors state that when they added 11 tons of balls to the 45 pct volume ball load, the power needle did not budge. The question arises as to whether any increase in capacity accompanied the 11-ton addition. It is hardly conceivable that all of the balls making up the 11-ton addition were sterile. However, as the work done in the ball mill is not by kilo-watthours per se or balls per se but by kilowatthour-balls, it is possible that the kilowatthourballs in the mill before the 11-ton addition had a different amperage, voltage or robustage than those in the mill after the addition and that the overall outcome was that equal work was done before and after the addition. The authors can keep the efficiency that they got with the 45 pct volume 1-in. ball charge and get additional capacity by putting more kilowatthourballs to work in the form of the addition of 11 tons of 2-in. balls to the present ball charge, and subsequent makeup additions by weight of 20 pct 2-in. balls and 80 pct 1-in. balls. J. F. Myers (authors' reply)-—Mr. Garms points his finger at the 11 tons of "sterile" balls added to the Tri-cone mill. A fact so interesting to us that we felt it should be reported. Clearly, ball slippage at the existing mill speed prevented any further transfer of power from the shell to the ball mass. We agree that the kilowatthourballs would increase, were we charging rough 2-in. balls as suggested by Mr. Garms. The assumption is, of course, that we would get a corresponding increase in work done. Our reported data (CIME meeting, Vancouver, 1946) shows a gain of work accomplished of 7.6 pct as we decreased the ball size from 2 in. to 11/2 in. and a further gain of work accomplished of 2.1 pct when we went from 11/2 in. to 1 in. The small ball results reported by H. R. Banks at Chapman Camp are also very convincing. In the case of 2-in. balls they would take the conventional ball path and start cascading. They would thus destroy the classifying pool over the "foot" of the

Jan 1, 1951

By J. F. Myers, F. M. Lewis

W. I. Garms-—The authors state that when they added 11 tons of balls to the 45 pct volume ball load, the power needle did not budge. The question arises as to whether any increase in capacity accompanied the 11-ton addition. It is hardly conceivable that all of the balls making up the 11-ton addition were sterile. However, as the work done in the ball mill is not by kilo-watthours per se or balls per se but by kilowatthour-balls, it is possible that the kilowatthourballs in the mill before the 11-ton addition had a different amperage, voltage or robustage than those in the mill after the addition and that the overall outcome was that equal work was done before and after the addition. The authors can keep the efficiency that they got with the 45 pct volume 1-in. ball charge and get additional capacity by putting more kilowatthourballs to work in the form of the addition of 11 tons of 2-in. balls to the present ball charge, and subsequent makeup additions by weight of 20 pct 2-in. balls and 80 pct 1-in. balls. J. F. Myers (authors' reply)-—Mr. Garms points his finger at the 11 tons of "sterile" balls added to the Tri-cone mill. A fact so interesting to us that we felt it should be reported. Clearly, ball slippage at the existing mill speed prevented any further transfer of power from the shell to the ball mass. We agree that the kilowatthourballs would increase, were we charging rough 2-in. balls as suggested by Mr. Garms. The assumption is, of course, that we would get a corresponding increase in work done. Our reported data (CIME meeting, Vancouver, 1946) shows a gain of work accomplished of 7.6 pct as we decreased the ball size from 2 in. to 11/2 in. and a further gain of work accomplished of 2.1 pct when we went from 11/2 in. to 1 in. The small ball results reported by H. R. Banks at Chapman Camp are also very convincing. In the case of 2-in. balls they would take the conventional ball path and start cascading. They would thus destroy the classifying pool over the "foot" of the

Jan 1, 1951

By Howard N. Eavenson

WASTE of coal, or perhaps more properly the percentage of its recovery in mining, has keenly interested me during an experience of over a half century in coal mining. In the early part of that time and until a few ears after World War I the use of better mining plans and supervision resulted in an increased recovery. Since then, and particularly since controlled prices have been effective and mechanized mining has been extensively practiced, the tendency has all been toward lesser recovery in mining. The purpose of this paper is not to criticize the methods being used, which after all are the result of economic forces, nor to point to any particular mining method or to any locality where the best recovery is obtained. but to call attention to the tremendous waste of our national coal resources that is constantly continuing. A possible method of reducing this loss will be suggested. but without any illusions that it will be followed or that much attention will be paid to it.

Jan 1, 1946

By P. W. Neurath, R. E. Waite

INFORMATION is available on the effect of stress on permeability, coercive force, and hysteresis for a variety of materials' but for the technically important grain-oriented 3 pct Si-Fe' the stress effects are not adequately covered in the literature; in addition, most of the stress data do not permit estimation of what would happen to the 60 cycle watt losses of the material under stress. Direct measurement of these effects therefore has been made and appears to be of considerable theoretical as well as practical interest. Losses were measured on the central 25 cm portion of 40~3~0.035 cm samples which could be stressed in a small tensile machine built for this purpose. Under tension the following effects, illustrated in the figures by a particular example, were observed: 1—"Unstrained" material of commercially acceptable loss level shows no change or a small increase in loss at low elastic stresses. For the sample shown in the figures, this is illustrated during the initial run, where a 7 pct increase is observed (Fig. 1). 2—There is no additional change in loss when the sample is further strained up to the elastic limit (Fig. 1). Within that limit the loss-stress curve is reversible. This is not illustrated in Fig. 1. 3—Even in the plastic range only small increases in loss are observed as long as the sample is kept under stress (Fig. 2, point N); often no increase can be observed (Fig. 1). 4—A large increase in watt losses is observed after removal of the stress subsequent to plastic deformation (Figs. 1 and 2). 5—All, or nearly all, of the foregoing increase in loss due to plastic deformation can be eliminated by a reapplication of tensile stress but returns upon removal of this stress. This is illustrated in the third run (Fig. 3). This run remains within the elastic limit and shows the reversibility of the stress-loss curve under this condition. These findings have been confirmed on large Samples (approximately 200~10~0.035 cm and larger) except that under 1 no initial increase in loss was ever observed on these. The accuracy of the watt loss measurements is of the order of 1 pct. Neither bending of the sample in the tensioning device nor a redistribution of flux along the width of the sample accounts for these results. They therefore must be explained on the basis of either a change in the domain configuration due to stress or a change in the domain wall mobility. Acknowledgment The assistance of Mr. Robert L. TenBroeck with the instrumentation is gratefully acknowledged.

Jan 1, 1956

By F. D. Rosi, C. A. Dube, B. H. Alexander

The slip and twinning planes have been determined in deformed crystalsof titanium by an X-ray method of analysis. The slip planes are of the type {1010} and {1011}, while the twinning planes are of the type {1072}, {1121}, and (1122). In the case of the predominant (1010) slip, a type I digonal axis of indices <1120> was the effective slip direction. Microscopic observations on the appearance of the three twin types disclosed a distinct difference in their shapes, and an attempt was made to correlate this difference with the amount of twinning shear. The slip mechanism was discussed on the basis of the difference in c/a ratios for the common hexagonal close-packed metals. IT is a general rule that slip in metals occurs most easily on the planes of greatest atomic density and of the largest interplanar spacing, and that the direction of slip is a close-packed direction. Thus, of the hexagonal close-packed metals which have been investigated by the German school in the decade 1925 to 1935 (cadmium, zinc, and magnesium), it has been found that slip occurs on the basal plane (0001) and in a <1120> direction.&apos; In the case of magnesium above 225oC, slip occurs on a first-order pyramidal plane (1011) as well as on the basal plane.&apos; The hexagonal metals, moreover, deform in part by twinning, and, again with the exception of magnesium at elevated temperatures, the twinning plane has always been reported to be a second-order pyramidal plane {1012).&apos; In view of this consistency in the plastic behavior of these metals, it has been supposed that the same slip and twinning elements would be operative in the other important hexagonal close-packed metals (titanium, zirconium, and beryllium), although there are some reasons why this may not be so. Referring to the tabulation in Table I of the axial ratios of the important metals in this structural class, it may be seen that they can be categorized into three distinct groups depending on the value of their c/a ratios. In the first group are cadmium and zinc with large positive deviations from the ideal ratio for closest packing. This expansion in the direction of the c-axis would be such as to accentuate the basal plane as the slip plane, because it increases the interplanar spacing. The second group includes cobalt and magnesium, which closely approximate the ideal ratio. In the third group are zirconium, titanium, and beryllium whose axial ratios are decidedly less than the ideal. The lattice of the elements in this group is compressed along the c-axis which, in effect, tends to make the basal plane less favorable for slip inasmuch as it reduces the interplanar spacing and the atomic density of this plane. Thus, for metals in the third group, it may be said that there is a certain indefiniteness about the glide plane, in that planes other than the basal plane may be expected to operate. The fact that magnesium begins to exhibit planes other than the normal slip and twinning planes at elevated temperatures may indicate that this behavior will become more pronounced in metals of lower axial ratios. For this reason it might be expected that the slip and twinning behavior of zirconium, titanium, and beryllium would be different from that found in cadmium and zinc. One indication that there is a difference is reflected in the recent work which has been done on the deformation textures of these metals."&apos; As will be shown later, there is a significant difference between the rolling textures of cadmium, zinc, and magnesium on the one hand, and zirconium, titanium, and beryllium on the other, and this difference suggests that the mechanism of plastic flow is not the same in these two groups of metals. From these considerations, therefore, it may be seen that there are good reasons why it cannot be concluded that titanium will have the same crystal-lographic elements of slip and twinning as those exhibited by zinc and cadmium. For the complete determination of these elements it would be most advantageous to have single crystals. However, much valuable information can be obtained by using coarse grained polycrystalline specimens, provided the grains are sufficiently large to permit X-ray

Jan 1, 1954

By R. Norman Orava

The strain-mte behavior of. the tensile yielding and flow characteristics of arc-cast polycrystal-line molyhdenum of pidrity > 99.97 pct was investigated in the range 10-6 to 10-1 sec-1 at 243o, 293o, and 373oK for two grain sizes, 0.027 and 9.098 mm. For 0.998 mm, the strain-mte dependence of the lower yield stress can be described by a straight line of the jowl a, = a + 0 log e over the whole range of strain rates examined. However, for 0.027 mm, the data fit several consecutive linear relationships of the above form. The previous association of these different linear regions with different rate -controlling mechanisms is questionable for molyhdenzdm. The senlilog strain-rate dependence of the yield stress is consistent with Hahn&apos;s model for yielding when the stress dependence of the mobile dislocation density, pr, is taken into account. Since p&apos;, from rate -cycling experirnents, is found to be relatively insensitive to stress for single crystaIs, the obserzied log-log dependence of their yield stresses is also in agreement with the model. A correction for the stress dependence of p&apos; changes the activation oolurrze for yielding from a zero to a more realislic inverse function of stress where the Peierls mechanism is believed to be rate-controlling. The gmin-size dependence of the lower yield stress is not a monotonic junction of strain rate. RECENT experimental evidence&apos; has revealed that the strain-rate dependence of the tensile lower yield stress, a,, of commercially pure ni- obium and tantalum tested at room temperature can be described by several consecutive semilog relations of the form: sv =a + ß log € [1] where i is the applied strain rate and a and 0 are constant Over only relatively small ranges of strain rate. Three linear regions were found for niobium and four for tantalum. Each linear region was interpreted as being associated with a different rate-controlling mechanism on the basis of the current

Jan 1, 1964

[MINING METHODS PHILIP B. BUCKY, Chairman WILLIAM J. COULTER. Past Chairman J. B. HAPPER, Vice-Chairman LELAND A. WALKER, Vice-Chairman, CLYDE E. WEED, Vice-Chairman E. D. GARDNER, Secretary]

Jan 1, 1946

By T. E. Seidel, A. U. MacRae

The dependence of the electrical and crystalline properties of silicon containing ion implanted boron atoms have been studied as a function of the incident dose, substrate temperature, and annealing tempera-ture. Hall effect studies show that the compensation due to defects is negligible after annealing at and above 500°C for a dose of 1013 ions per sq cm. Steps occur in the isochronal annealing curve for the sheet resistance at 300°C and 800°C and the mechanisms responsible for these steps are discussed. The range of the ions at 150 kev is 3500 * 5001. Electron diffraction studies on samples implanted with 150 ker ions show the existence of damaged (polycrystalline) silicon closer to the surface than the range of the ions. Etching and subsequent annealing of etched and unetched specimens suggest that the poly crystalline silicon is closer to the surface than the bulk of the distribution of ions. THE purpose of this study is to investigate some electrical and crystalline properties of ion implanted boron in silicon. Sheet resistance, Hall measurement, depth distribution by chemical staining, thermal probe and electron diffraction have been used to study the properties of samples implanted in a random direction with ions having energy up to 150 kev. The boron-silicon system has been the subject of considerable study&apos;-7 due to its possible application in the fabrication of electronic devices. The implanted ions produce considerable crystalline damage to the substrate crystal before they lose most of their primary energy and come to rest. Fortunately, most, if not all, of this damage can be removed by subsequent annealing or by holding the substrate at an elevated temperature during the implantation process. Sheet resistivity measurements of the p-type layers produced by boron implantation into room-temperature silicon, reveal that the annealing occurs in a two-step process: the first occurring at -300°C and the second at -800°C.&apos; Electron-microscope and low-angle electron diffraction studies of this damaged material indicate that the bombarding ions produce an amorphous silicon region about the trajectory of each ion.&apos; At high doses, these damaged regions overlap and an amorphous surface layer results. Furthermore, the 800°C annealing step, appears to be associated with the elimination of this gross damage. The depth distribution of the as-implanted and also the annealed samples have been studied by an angle lap and staining technique as well as by Hall measurements.3-7 We find the experimentally determined ranges of the boron are less than those calculated by the Lindhard-Scharff-Schiott theorys (hereafter referred to as the LSS theory). Our results are in essential agreement with pre- vious results. In addition, we have made Hall measurements on samples bombarded with a low dose (2 x 1013 ions per sq cm) to minimize impurity banding effects, and find that deep ionized levels influence the carrier concentration and room-temperature hole mobility until the sample is annealed at temperatures above 500°C. Above this temperature, the data corresponds to that obtained when silicon is doped with boron by conventional techniques. We have also made a depth distribution determination of the crystalline damage due to the implanted boron by using glancing angle electron diffraction techniques. It was found that the gross damage depth distribution of heavily implanted samples is closer to the surface than the depth distribution of the electrically active boron. By removing this gross damage, without affecting the distribution of the implanted boron, we have determined that the 800°C annealing stage is not exclusively due to the improvement in the crystalline perfection of the amorphous material but also appears to be associated with the annealing of some point defects. These results and their relation to previous results are discussed in the following sections. IMPLANTATIONS Boron ions were implanted into silicon samples using an accelerator capable of operating from 10 to 150

Jan 1, 1970

By B. H. Caudle, L. H. Silberberg

Reservoir depletion by natural water drive is typified by the movement of water from an aquifer into the adjacent oil-bearing formation. Prior studies of this type 01 water movement have generally neglected the resistance to flow in the aquifer and its effect on the movement of water into the oil bearing zone. A method for designing and operating scaled models of such reservoir systems is presented. Experimental data on a model of an edge-water-drive reservoir are shown and discussed. INTRODUCTION Most petroleum reservoirs derive at least a part of their productive capacity from water influx. This water may be injected from the surface or it may come from an aquifer adjacent to the oil zone. In either case, the reservoir engineer must be able to estimate the advance of the water as a function of either the elapsed time or the fluid produced. This paper describes a type of fluid flow model which includes the effect of viscous and gravitational forces in the reservoir and the surrounding aquifer. Natural water influx can be divided into three general (but widely overlapping) categories according to the direction of flow in the aquifer. These are shown schematically in Fig. 1. Fig. l(a) illustrates an edge-water-drive mechanism. In this case, water advances updip along the stratum but the movement of water is essentially horizontal and very little of the oil is actually underlain by water. The bottom-water drive (Fig. l(b)) is characterized by a thick aquifer underlying the oil zone. The water movement is generally vertical in the aquifer. The third category is illustrated in l(c). Referred to as the "thin oil column", this type of oil accumulation consists of oil over water in a relatively thin stratum. Fluid movement is horizontal in both the oil and water zones, except close to the producing well. One characteristic which all three of these types have in common is that part of the water flow takes place in the water invaded region of the oil reservoir while the remainder of the water flow occurs in the aquifer. The resistance to the flow of water usually will not be the same in these two regions. This occurs because the porous rock normally contains only water in the aquifer, while the water-invaded region always contains microscopic globules of bypassed oil (residual oil) which interfere with the flow of water through the rock. The production history of any type of a water-drive system is a function of two phenomena — where the water goes, and how it displaces the oil in the area invaded by the water. It is usually impractical to study these two phenomena at the same time. In most cases one type of analysis is used to predict the gross movement of water in the reservoir, while a different type of analysis is employed to determine the amount of oil to be recovered from the gross volume contacted by the water. The portion of the oil reservoir which is invaded by water is mainly a function of the resistances to fluid flow in the several parts of the flow system. This gross water influx in the three types of natural water drive typically results in water cusping to the production wells either from the side or from below (water coning). In 1747, Muskat described a greatly simplified mathematical model for predicting the area invaded by water in a bottom-water-drive reservoir.1 This model used the Laplace equation and suitable boundary conditions to describe the isopotentials and streamlines in the flow system. The use of this model assumes that (1) the reservoir rock is homogeneous in nature, (2) the oil and water mobilities are equal, (3) there is no oil flowing in the water invaded region, and (4) external forces — such as gravity — do not affect the flow. This method has been used for both natural water drives and water injection projects. In this, as in most models, the method of images is used to reduce the size of the model necessary to describe the reservoir. The assumption of equal viscosities and permea-

Jan 1, 1966

By A. L., Bailey

All who are even generally aware of the tremendous rate of increase in coal washing operations must realize the growing importance of the float-and-sink test. I believe it is conservative to estimate that, in the past decade, the dollar volume of float-and-sink testing has increased tenfold. It is a simple matter of economy, then, to examine the factors that determine the cost of adequate float-and-sink testing. When the Coal Preparation Section of the Bureau of Mines entered upon a greatly expanded program of such work in connection with the synthetic liquid fuels investigations, it seemed advisable to examine these factors experimentally. The principal consideration that differentiates float-and-sink test sampling from general purpose sampling, is that the original particle size must be preserved. Therefore, the total cost of the test will be directly affected by any standard that might be proposed to limit sample bulk reduction at any given particle size. For this reason, the relationship of sample size to variability of results was the first factor to be studied experimentally.

Jan 1, 1949

By C. A. Stickels, C. M. Yen

Secondary recrystallization was investigated in vacuum-melted electrolytic iron to which 70 pm N was vacuum-meltedadded. The secondary texture is "near {554}<225>" for material cold-rolled 75 to 90 pct, the sharpness of the texture increasing with increased rolling reduction and with decreased annealing temperature. At reductions of 95 and 97.5 pct the secondary texture is &apos;"near {322)(296)". Both secondary orientations also exist as major components of the primary re-crystallization texture. Development of a strong "near {554) (225)" secondary texture appears to depend on the evolution of the Primary texture to a transition texture depleted in orientations near the secondary orientation before the onset of secondary growth. A variety of qualitative experinzents have been used to show that nitrogen is important in limiling primary grain growth. The presence of nitrogen does not seem essential for the establishment of a transition texture, but a loss of nitrogen during annealing may facilitate growth of grains in the secondary orientation. Secondary grains we shown to form initially at the specimen surface. This is not thought to indicate that surface energies are important in the growth process. It is proposed that the quasi-two-dimensional character of surface grains permits discontinuous growth parallel to the surface before secondary growth of interior grains is possible. An earlier study of recrystallization textures in 90 pct cold-rolled electrolytic iron showed that secondary recrystallization occurred after annealing for several days at 700C1 This type of secondary recrystallization, which had not been reported previously, results in the formation of a strong texture, best described by the indices "near {554}(225)". The purpose of the present work was to investigate the effect of various processing variables on secondary recrystallization in this material and determine the mechanism of secondary grain growth. LITERATURE REVIEW An understanding of the mechanism of a secondary recrystallization process depends on knowing: 1) how grains in the secondary orientation come to be in the primary recrystallization texture; 2) why normal grain growth does not occur; and 3) what factors determine the strength of the secondary texture. For secondary growth of grains of a particular orientation, a certain minimum fraction of the grains must be in that orientation after primary recrystallization. This requirement is apparently satisfied "naturally" in certain systems, i.e., when the primary texture obtained by rolling and recrystallizing material initially randomly oriented contains a sufficient fraction of primaries in the secondary orientation. However, in other cases, e.g., {110}<001> and {100}<001> secondary growth in silicon iron,2 it is necessary to enhance the fraction of primary grains in the secondary orientation by rolling and recrystallizing textured material. In the present case, the "near {554}<225>" orientation is contained within the spread of orientations found in the primary recrystallization texture of iron or bbc iron-base alloys. In systems where the main driving force for secondary growth is the reduction in total grain boundary energy, secondary growth is observed only when normal grain growth is minimized. Four ways in which normal grain growth can be limited are: 1) Limitation by a strong primary texture. When a very strong primary texture consisting of a single component or twin-related components develop, most primary grains are separated from one another by relatively immobile small-angle grain boundaries. The classic instance of this is secondary growth into the primary cube texture in some fcc metals. 2) Limitation by precipitates. Precipitates present in the proper volume fraction with a suitable dispersion will limit primary grain growth. The role of MnS inclusions in impeding normal grain growth in Si-Fe is well-documented.5 3) Limitation by sheet thickness. Normal grain growth slows drastically when the mean grain diameter is of the same order as the sheet thickness. This effect has been used to obtain secondary recrystallization in thin sheets of high-purity silicon iron.&apos; 4) Limitation by solute impurities. It is well-established that certain impurity elements in solution can have a large effect on grain boundary mobility.&apos; However, there does not seem to be any secondary recrystallization process in which primary grain size stabilization has been shown to be due to the drag exerted on grain boundaries by dissolved impurities. In certain systems, e.g., secondary recrystallization in silver,&apos; the means by which normal grain growth is limited has not been identified, and solute-impurity limitation might be suspected. In order to understand the factors which determine secondary texture strength in three-dimensional growth, it is necessary to examine in more detail the current picture of general secondary recrystallization processes. Following Cahn,9 it is assumed that the primary grains have a range of sizes and that secondary growth of one of the large grains in this distribution is possible when it exceeds a critical size with respect to its neighboring grains. The critical size depends on the ratio ?S/?p, where ?s is some sort of average grain boundary energy between the potential secondary and the primary grains and ?p is some sort of average grain boundary energy between primary grains. For a constant primary grain size, the critical size for secondary growth increases as ?$/?p increases. May and Turnbull5 have incorporated the

Jan 1, 1969

By Charles C. Peterson, Ronald E. Enstrom

Significant improvements have been made in the ursine systern for epitaxial vapor gvowtlz of Gds. The electron concentration has been reduced to below 1015 cm-3 with electron-mobility values as high as 7200 sq on per v-sec at 30PK and 60,000 sq cw7 per v-sec at 77°K. These high-purity layers have been successfully gvown on n&apos; layers with thicknesses as snzall as 1 p and as large as 125 p. Furthermore, high-purity layers have been incorporated into an structure without degradation of electrical properties. Different impurity gradients at the n&apos; to n interface were achieved. Using these structures, Gunn oscillators have been operated at frequencies as low as 0.4 GHz and as high as 40 GHz, which equals the highest valzle yet reported. The frequencies nleasured are in reasonable apeenlent with those calculated fron7 the n-layer widths and the transit times. CW operatiolz has been achieved from 3 to 20 GHz. In Gunn devices for >40 GHz operation (i.e., witk an n-layer tkickness below 2.5p), the n layer has been generally too pure to permit oscillations to occuv. ThE mechanism of the Gunn oscillator&apos; is now generally understood.2-5 Electron transfer from the (000) minimum to the ( 100) minima leads to a voltage-controlled negative resistance, which results in the formation of domains of high and low electric fields which travel through the crystal with a characteristic velocity of about 107 cm per sec. When a constant voltage is applied to the sample, this domain motion generaliy leads to oscillations at the transit time frequency, fT = 107/L, where L is the thickness of the active n layer.&apos; In this case, a single high-field domain is nucleated at the cathode and grows until it reaches the anode, where it is absorbed; a new domain then forms at the cathode.5 There are related modes of oscillation of the same device in which a tuned circuit causes a periodic variation of the voltage across the device. The frequency then may be either above or below fT. It has been shown that oscillations will occur when n • L 2 1012 cm"2, where n is the carrier concentration in the active region. At lower doping densities, 10" 5 n . L 5 lo&apos;&apos;, the doping is said to be subcritical for oscillation and the description of the device operation changes since the domains become comparable in thickness to the device thickness, L. Nevertheless, such devices have been found to be "active" and, in fact, are useful as amplifiers.778 For CW operation, one wishes to minimize the power dissipation, which is related to the electric field, E, and the low-field resistivity, p. Below threshold, the power dissipation density is approximately Typically, at threshold where E -- 3 x 103 v per cm, the dissipation density is 107/p watts per cu cm, p in ohm-cm. Furthermore, it is desirable to provide good conduction of the heat generated in the active region. The temperature rise, AT, across the active layer of a sandwich structure, with a heat sink on one side, is given by AT = (E&apos;/P)(L~/~~) where it is assumed that the thermal conductivity, u, and resistivity, p, of GaAs do not vary appreciably over the range of temperature and electric fields existing in the layer. This expression shows that, subject to maintaining an appropriate (n . L) product, CW operation is favored by a higher resistivity, a lower threshold field, and a thinner device. The above device requirements place some severe restrictions on the materials to be used, especially for high-frequency (>10 GHz) operation. Here, it is necessary to prepare very thin (<10 µ) layers of GaAs, in a state of high purity (-1 ohm-cm), complete with very low-resistance ohmic contacts. Problems of handling such thin and fragile samples become inereasingly severe as the frequency is raised. Furthermore, while good low-resistance ohmic contacts are easily applied to low-resistivity GaAs, contacting difficulties are encountered for these higher-resistivity materials.

Jan 1, 1968

By John Parker

During 1965, a new stoping method was developed at the Bunker Hill Mine at Kellogg, Idaho to enable the profitable mining of a large low-grade zinc orebody. This method, called the Bunker Hill pillar system, incorporates the use of the air-powered Atlas Copco T2GH or CAVO 310 as the prime production component. The method and equipment proved highly successful and was the first step in mechanizing what had been a classic example of a square-set and horizontal cut-and-fill operation. Since that time there has been a realization that the small rubber-tired type of loader could be effectively used to advantage in increasing productivity and decreasing costs in many different stoping and development applications throughout the mine. The net result since 1967 has been that a large and increasing portion of the mine production and an increasing share of the development is being done with small rubber-tired loaders. Presently, ten Atlas Copco T2GH and six CAVO 310 air-operated loaders are in use, having replaced the time-honored slusher which was first introduced into the Bunker Hill Mine in about 1898 in the form of a slip scraper. The recent advent of diesel powered loaders in the one yard size range has in 1970 further accelerated the trend toward rubber-tired mobile equipment at the Bunker Hill Mine. Although The Bunker Hill Company is using small diesel loaders of several different types and manufactures for inclines, declines and level development and will undoubtedly find stoping applications for this equipment, .the bulk of this paper will be confined to the use of the air-operated loader, leaving the diesel equipment to a future presentation. It is sufficient to say that we feel at this time that the Autoloader type air-powered equipment and the small diesel front end-loader

Jan 1, 1970

By Matthew J. Hrebar

SINCE the first installation in 1936, the use of rubber-tired equipment for secondary mine haulage has increased rapidly. From August 1936 to January 1944, approximately 750 rubber-tired units were placed in operation. Of these, 650 units were installed after January 1940. During the year 1941, it per cent of the mobile loaded bituminous tonnage was hauled by rubber-tired equipment. This percentage increased to about 20 per cent during 1943. The first rubber-tired units were battery-powered tractor-trailer types. In the year 1938 the battery-powered self-contained car was introduced. In 1939, the cable-powered self-contained car made its appearance. These self-contained units are our modern "shuttle cars." By the end of 1941, approximately 15 per cent of the shuttle cars in operation were cable-powered. Installations made during 1942 and 1943 increased the proportion of cable-powered cars to about 40 per cent of the total number of shuttle cars in use. There is a marked trend toward cable-powered cars, although one need not make many trips to mines where both types are used to discover that opinions are highly controversial as to the relative advantages and field of use of each type. It is natural for mining engineers in the coal-mining industry to have diversified opinions regarding the use and application of coal-mining machinery. Coal mining is not a standardized industry; natural conditions vary from mine to mine. However, it is surprising how many operations in scattered fields arrive at the same conclusions and adopt similar methods. This happens not so much through cooperation and exchange of ideas as through hard, costly experimentation, involving months of low production and relatively high costs. What then, are the reasons for increased use of cable cars? Could it be that present scarcity of materials is forcing their use? True, battery cars require considerably more auxiliary equipment, but during 1942 it was no more difficult to obtain than shuttle-car cable. Furthermore, the trend toward cable cars started before curtailments on mining equipment took effect. Undoubtedly, cable-reel shuttle cars were introduced to produce coal at a lower cost by simplifying trackless mining through the elimination of auxiliary equipment, and by reducing the first cost on equipment. The theory of cable-powered shuttle cars is well based, as can be proved by successful operations. Many operators have weighed the advantages and disadvantages of each type carefully before choosing the type most suitable to their operation. Others have taken much for granted and accepted one or the other without much thought. At many small operations, low first cost is apt to be the deciding factor. One large company, with considerable experience in the use of battery locomotives, preferred to use cable cars even though it was neces¬sary to reduce the section voltage from

Jan 1, 1944

By L. G. Evans, W. W. Simpson, W. A. McKinney

In conventional dump leaching of strip wastes from open-pit porphyry mining operations, many years are required to extract a fraction of the copper from the contained sulfide minerals. Furthermore, none of the molybdenum, gold, or silver is recovered by current leaching practices. Because the copper sulfide minerals and molybdenite generally are distributed in the minute fractures throughout the host rock, the mineral values tend to concentrate in the finer fraction of the strip material. This suggests the possibility of segregating the enriched fines in as-produced or coarsely crushed strip wastes for blending with the ore grade flotation mill feed. In the past, extra handling or crushing of mine wastes to enable better leach extraction has been deemed uneconomical. However, recent articles on mobile crushing plants and conveyor haulage systems indicate possible cost improvements for these techniques over conventional truck haulage of ore and mine wastes. Another potential benefit from removal of the fines from copper-bearing mine waste derives from the increased permeability of the coarse dump material to leach solutions and air. Better permeability should result in faster oxidation of the sulfide minerals, less channeling of leach solutions, and fewer perched solution tables caused by compacted fines. Therefore, in addition to the metal values recovered by flotation, higher copper recoveries by leaching also might be achieved. The Salt Lake City Metallurgy Research Center of the U.S. Bureau of Mines undertook an investigation to determine the feasibility of removing enriched fines from strip waste material for separate treatment by froth flotation for early recovery of part of the copper, molybdenum, gold, and silver. Percolation leaching tests were made on mine-run and sized fractions of the dump material to determine the effect of fines removal on copper extraction. Samples Tested Three 10-ton samples of copper sulfide-bearing mine waste were obtained for the investigation. Partial chemical analyses of the samples are given in [Table 1] and the consolidated results of detailed screen analyses made on the samples are given in [Table 2]. The -1/4-in, fine fractions contain from 16 to 32% of the copper and 19 to 40% of the molybdenum in 11 to 18% of the total waste. The samples also were crushed to -3 in. to determine the feasibility of crushing to obtain a higher distribution of the metal values in an enriched fine fraction while providing a more leachable coarse fraction that could be transported by a conveyor system to a dump. The -%-in. fine fractions screened from the crushed mine waste were almost identical in grade to the natural fines. From 2 to 4% more1/4-in. fines were produced by crushing to -3 in. Copper distribution was increased by about 3% over that in the natural fines, while molybdenum distribution was increased by about 10%. The -1/4-in. fines screened from the mine-run and crushed strip waste were almost as high in grade as the ore being treated at the mills. In addition, the fines

Jan 1, 1974

By C. S. Land

Relative permeability functions are developed for both two- and three-phase systems with the saturation changes in the imbibition direction. An empirical relation between residual nonwetting-phase saturation after water imbibition and initial nonwetting-phase saturations is found from published data. From this empirical relation, expressions are obtained for trapped and mobile nonwetting-phase saturations which are used in connection with established theory relating relative permeability to pore-size distribution. The resulting equations yield relative permeability as a function of saturation having characteristics believed to be representative of real systems. The relative permeability of water - wet rocks for both two- and three - phase systems, with the saturation change in the imbibition direction, may be obtained by this method after properly selecting two rock properties: the residual nonwetting-phase saturation after the complete imbibition cycle, and the capillary pressure curve. INTRODUCTION Relative permeability is a function of saturation history as well as of saturation. This fact was first pointed out for two-phase flow by Geffen et al. 1 and by Osaba et a1.2 Hysteresis in the relative permeability-saturation relation also has been reported for three-phase systems.3 Since saturations may change simultaneously in two directions in a three-phase system, four possible relationships arise between relative permeability and saturation for a water-wet system. The four saturation histories of this system were given by snel14 as 11, ID, DI and DD. I and D refer to the direction of saturation change (imbibition and drainage), with the first letter of the symbol indicating the direction of change of the water phase. As used in this paper, the second letter of the symbol refers to the direction of saturation change of the gas phase, i.e., D and I indicate an increase and decrease, respectively, in gas saturation. Only a few three-phase relative permeability curves have been published. Leverett and Lewis5 published three-phase curves for unconsolidated sand, and Snell4 reported results of several English authors for both drairrage and imbibition three-phase relative permeability of unconsolidated sands. Three-phase relative permeability curves for a consolidated sand were published by Caudle et a1.3 for increasing water and gas saturations (ID). Corey et a1.6 reported drainage (DD) three-phase relative permeability for consolidated sands. Recently, Donaldson and Dean7 and Sarem8 calculated three-phase relative permeability curves from displacement data on consolidated sands, also for saturation changes in the drainage direction. The only published three-phase relative permeability curves for consolidated sands with saturation changes in the imbibition direction (11) are those of Naar and Wygal.9 These curves are based on a theoretical study of the model of Wyllie and Gardner 10 as modified by Naar and Henderson.11 Interest in three-phase relative permeability has increased recently due to the introduction of new recovery methods and refinements in calculation procedures brought about by the use of large-scale digital computers. The scarcity of empirical relations for three-phase flow, and the experimental difficulty encountered in obtaining such data, have made the theoretical approach to this problem attractive. RELATIVE PERMEABILITY AS A FUNCTION OF PORE-SIZE DISTRIBUTION Purcell used pore sizes obtained from mercury-injection capillary pressure data to calculate the permeability of porous solids.12 Burdine extended the theory by developing a relative permeability-pore size distribution relation containing the correct tortuosity term.13 The works of Purcell and Burdine were combined by Corey14 into a form that has

Jan 1, 1969

By J. E. 214-000-000-010 Warren, M. R. J. Wyllie, T. Meidav, L. Scharon, R. Uhley, A. J. deWitte

By considering the stoichiometry of the underground combustion process, an equrztion was derived relating the point velocity of the combustion front as a function of the air flux, fuel content, efficiency&apos; of oxygen uitilization, hydrogen:carbon ratio, and ratio of carbon dioxide In carbon monoxide produced. In addition, an equation for cnlcrllating the amount of injected air necessary 10 produce the oil from I Cu. ft of reservoir space was derived. This equation has a direct bearing on evaluating compression costs for the process. Values calculated from these equations are in good agreement with experimentnlly obscrvcd values obtained from laborator) tube runs reported in the literature. The equations are also u\eful for atlcrlyzinf: experinlental data for inconsistencie.r.. I NTRODUCTION The injection of energy in the form of heat into an oil-bearing formation as a means of increasing the recovery of oil has been the subject of much discussion and experimentation as far back as 1920. The ideas have ranged from the injection of steam to the direct use of electricity as a means of getting thermal energy into a formation. However, most laboratory and field studies have been confined to the process whereby a portion of the crude in the reservoir is burned to generate the necessary heat to increase the mobility of the oil by reducing its viscosity. A recent paper by W. L. Martin, et al.&apos;, described the results of a number of laboratory experiments in which the combustion process was studied in a sandpack. Studies such as this indicate that with the temperatures involved, the oil is partially coked in this process and it is this coke which provides the fuel for combustion. The combustion of this residual coke can be analyzed theoretically by applying simple stoich-iometric principles to the combustion process. It is the purpose of this paper to compare the results reported in these experiments with those obtained from theoretical considerations. The variables studied by Martin, et al, among others, include combustion front velocity, air flux, oxygen utilization, fuel concentration, ratio of carbon dioxide and carbon monoxide in combustion gases, and the hydrogen:carbon ratio of the residual fuel. All these variables come into consideration when the stoichiometry of the combustion process is studied. The experimental conditions and results are tabulated in Tables 3 and 4 of Ref. 1. For purposes of comparison with calculated results, it was necessary to change the units of a number of the variables. UNDERGROUND COMBUSTION STOICHIOMETRY The combustion or oxidation of the carbonaceous residue or "coke" on the sand can be described in terms of the equation: where n is atomic ratio of hydrogen to carbon, m is ratio of moles of CO2 to CO produced, and CH. is taken as the unit molecular weight, realizing that actually the coke has a large range of high molecular weights. Assuming that all the fuel laid down in the sand is burned, the ratio of moles of oxygen consumed to moles of coke can be written as follows. Moles 0MolesMolescoke In terms of moles O2 per pound of coke, where 12 + n is the molecular weight in pounds per mole. If the 0, consumed is measured in standard cubic feet (60°F and 1 atm) and the coke in pounds per cubic foot of formation, 2, then: