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By A. H. Daane, A. S. Wilson

The U-Cr system is of the simple eutectic type with some solid solubility of chromium in r and ß uranium. The eutectic occurs at 20 atomic pet Cr and melts at 859°C. The maximum solubility of chromium in y uranium is 4 atomic pet at the eutectic temperature, and in ß uranium the solubility is estimated to be 1 atomic pet. y-ß and ß-a transformations were found to occur at 737°estimatedt° and 612°C respectively. DURING 1944 and 1945, the U-Cr constitution diagram was studied in this laboratory as a part of a research program on uranium metallurgy in the Manhattan Project, and the work was described in a Manhattan Project report issued in December 1945. This paper is based on that report, which has been declassified. Prior to this study, it had been shown by other Manhattan Project workers that the low Cr-U alloys could be quenched to retain the form of uranium. Experimental The uranium used in this work was massive metal prepared in this laboratory and contained less than 0.1 pct of other elements. The chromium was 200 mesh powder obtained from the A. D. McKay Co. and was found on analysis to be 99.5 pct Cr with 0.3 pct Fe the major impurity. Alloys, weighing 400 to 600 g, were prepared by induction heating the components to 1700°C in slip-cast ZrO, crucibles in a vacuum of 3x103 mm Hg. TO prevent too violent agitation of the melt by the induction field with subsequent crucible breakage and sample loss, the ZrO2 crucible was placed in a graphite crucible, which was surrounded by a layer of powdered carbon insulation 2 to 3 cm thick. Polished vertical sections of the alloys were examined microscopically to confirm their homogeneity. Heating and cooling curves were taken on the alloys by reheating them in ZrO, crucibles to 1200°C and inserting a mullite-protected chromel-alumel thermocouple into the melt by means of a slip seal in the vacuum head of the furnace. A recording potentiometer traced the curves which had a slope of from 3" to 6" per min. Samples of the alloys were prepared for metallo-graphic examination by conventional mechanical polishing techniques followed by an electrolytic polish in an ethylene glycol-phosphoric acid-ethyl alcohol bath. The structure of the alloys was brought out clearly by this procedure so that no further etching was required. Samples for chemical analysis were taken from drillings from the top, center, and bottom sections of the alloys. The uranium was determined by titra-tion with Ce(SO1)2, while the chromium was titrated with FeSO,; the uranium and chromium totaled at least 99.6 pct in all of the alloys prepared. X-ray samples were prepared by filing bulk specimens in a helium-filled glove box and annealing the resulting powder in a zirconium-gettered helium atmosphere. A 114.6 mm diam Debye-Scherrer camera and a Weyland nonsymmetrical self-focusing camera were used with filtered copper radiation to obtain the powder X-ray diffraction data. Results The data obtained in this study have been combined to construct the constitution diagram of the U-Cr system shown in Fig. 1 where the arrests observed in cooling curves are indicated by dots. The liquidus arrest was quite distinct in thermal data taken on alloys in the range 0 to 20 atomic pct Cr. The eutectic arrest was not observed in studies on the 2.5 and 4.5 pct Cr samples but appeared in the 7.5 pct samples, which suggested some solubility of chromium in y uranium. On quenching from 859 °C, the 2.5 pct sample showed but one phase while the 4.5 pct sample contained a small amount of the eutectic along the grain boundaries; see Figs. 2 and 3. From this the maximum solubility of chromium in r uranium has been set at 4 pct. X-ray studies on these samples showed that the r phase was not retained at room temperature by quenching, but in each case a pattern was observed .which has been identified with the ß phase of uranium. Thermal data show the y-ß transformation of uranium lowered to 737°C as a consequence of this solubility. On quenching from the ß range (660°C), precipitation of chromium in the primary uranium is observed in the 2.5 and 4.5 pct Cr samples (see Figs. 4 and 5),

Jan 1, 1956

By J. P. Zannaras

Referring to the article by R. J. Charles and P. L. de Bruyn, let us assume that W = weight of glass bar; P = weight of hammer; e = total deformation; K = unit of deformation; K = potential stress energy; E = modulus of elasticity; L = length of the bar; 7 = coefficient of inertia; h = height, ft; V = velocity, fps; and a = cross section area. The portion of kinetic energy which is effective in producing stress energy in a fixed bar struck horizontally is given by the formula" 1 + 1/3 W/P K =1+1/3w/p/(1+1/2w/p)2 P.V2/2g = Ph where 1 + 1/3 W/P ? (1 + 3W/P/(1=1/2w/p)2 [8] Putting e e = W/P =------------ From the above equation it can be seen that the maximum transfer of kinetic energy to stress energy is when e = 0 or W/P = 0 which indicates that the weight of the hammer must be very large as compared with the weight of the impacted rod. Eq. 8 diametrically opposes the conclusions reached by the authors of this article. In fact, if their suggestions were followed to the extreme when e = co when P = 0, there would be no transfer of kinetic energy to stress energy at all, as 7, becomes zero. Eq. 8 presumes that the velocity with which the stress is propagated through the bar is infinite, whereas the authors claim that the compression waves reflected are reaching the struck end of the bar prior to the complete transfer of the kinetic energy to cause such modification of the conditions there as to make them reach the reverse conclusions demonstrated by the above formula. That such interference exists is unquestionably demonstrated by the authors and others. However, if my observations are correct, such interference for this specific experiment and also for practical comminu- tion is insignificant, and the conclusions of the authors are in error and must be reversed to comply with Eq. 8. Eq. 5, w. = AE 2/2, given by the authors on page 51, is derived from the following equation (Eq. 9): K = 1/2Pe, where P = Sa, S = ?E, e = EL, and L = 1. The above formula, Eq. 9, cannot be applied in this case. This formula is applicable for static loads where the load increases from zero up to its final value, P, in such a way that the deformation at different instants is proportional to the loads acting at those instants and actually represents the area of a right triangle in the strain load diagram of base e and height P. The typical photographs shown in Figs. 3 and 4 represent the familiar strain load diagrams, and since the line of the wave marks the existence and intensity of the strain with the unquestionable conclusion that such strain has been caused by the action of a load acting continuously all along the wave until it reached the horizontal axis, the work stored at this point is represented by the area under the wave line and the horizontal axis and not by the area of the fictitious triangle given by the authors. Then if this is correct, even visual estimation of these areas at gage stations given in the typical photographs of Figs. 3 and 4 suffice to contradict the authors' calculation given in Figs. 6a and 6b and Figs. 7a and 7b. The typical photographs presented by Charles and de Bruyn show a considerable variation of the intensity of the strain at different stations but very small variation of areas which actually represent the stress energy at the corresponding stations. And, apparently, by squaring the small quantities, the authors magnified their error tenfold. J. M. Frankland's paperV iscusses the relative strain intensity and not the total energy for different types of impact loading. He states in his paper, "The reader is explicitly warned not to confuse the results in this report with those obtained when the load is applied by a blow as from a hammer. In this case the peak load rises to very large but mostly unknown values. The accompanying large deflections and stresses are the result of high values of P, not of the dynamic load factor n. According to Frankland "the dynamic load factor" is the numerical maximum of the response factor. It therefore appears that the authors followed the same procedure in obtaining the relative strain energy ab-

Jan 1, 1957

By Bruce Chalmers, D. C. Larson

Aluminum crystals with longitudinal-axis orientations of (111) . (110), and (100) were deforined in tension and annealed. The conditions of deformation were controlled so that the re crystallization nuclei originated in either the heavily deformed regions at saw cuts {artificial nucleation) or in the lightly deformed matrix (spontaneous nucleation). The artificial-nucleatioln experiments showed that in lightly deformed (110) and (100) crystals low-angle twist boundaries are most mobile, while in (111> crystals and heavily deformed (110) and (100) crystals high-angle tilt boundaries with near (111) rotations are favored. The spontaneous-nucleation experiments showed the existence of preferred orientations in the (111) crystals. The nonrandomness of the grain orientations is quantitatively determined through a comparison with the results which would he obtained from a randowl set of grain ovientations. PREVIOUS recrystallization studies have been performed on single crystals deformed in tension.1 7 The crystals used in these studies usually had random tensile-axis orientations and the extent of deformation was not a primary consideration. The present study concerns the recrystallization of single crystals with tensile-axis orientations of (Ill), (110), and (100). The emphasis of this work is on the influence of the tensile-axis orientation and the degree of deformation on both the nucleation and growth processes. The multiple-slip orientations were chosen because secondary slip or slip intersection promotes nucleation.1,5,8 These crystals recrystallize at lower strains than the crystals which are oriented for single slip. Also, the greatest variation in deformation behavior is exhibited by the multiple-slip orientations. The stress-strain curves for crystals with tensile-axis orientations of (111) are higher than the stress-strain curves for poly-crystals, and the stress-strain curves for crystals with tensile-axis orientations of (100) are lower (at large strains) than the stress-strain curves for the crystals which deform initially in single slip.g The recrystallization nuclei originated in either 1) the homogeneously* deformed matrix of the crys- tals or 2) the heavily and inhomogeneously deformed regions at saw cuts. The nuclei will be referred to hereafter as spontaneous and artificial nuclei, respectively. The two terms do not imply a difference in the nature of the nuclei; they imply simply a difference in the mode of introduction of the nuclei. During spontaneous nucleation very few (always less than ten) grains nucleate, while during artificial nucleation large numbers of grains nucleate. Only a fraction of the artificially nucleated grains penetrate very far into the deformed matrix during annealing. The grains that penetrate the farthest into the deformed matrix will be referred to as the dominant grains. EXPERIMENTAL PROCEDURE The thirty-five crystals used in this investigation were grown from the melt in milled graphite boats at a rate of 1.6 cm per hr. The crystals had dimensions of approximately 6 by 12 by 80 or 6 by 6 by 80 mm and the aluminum was of 99.992 pet purity. The as-grown crystals were annealed at 610°C for 24 hr and furnace-cooled. They were then heavily etched and electropolished in a solution of five parts methanol to one part perchloric acid. The crystal orientations were obtained by back-reflection Laue photographs and were accurate to ±2 deg. The tensile-axis orientations were (loo), (110), and (111). Two of the side faces of the (111) crystals were (110) lanes. The (110) crystals had both {100) and {110) side faces and the (100) crystals had (100) side faces. The crystals were deformed at a strain rate of 0.003 per min. Shear stress and shear strain were obtained by multiplying and dividing the tensile stress and strain, respectively, by the Schmid factor, m. For the (111) crystals m = 0.272 and for the (110) and the (100) crystals m = 0.408. The Schmid factor is effectively constant during deformation for all orientations. The deformed crystals were sawed into 1-in.-long specimens while the crystals were totally enclosed in a graphite boat. The sawing was performed very carefully in order to limit the plastic deformation to the sawed regions. The specimens were electropolished in the solution mentioned above to remove the sawed-end deformation as well as controlled amounts of surface material. A special stainless-steel grip was used to hold the specimens during the electropolishing treatment. The gripping faces were flat, with no teeth, to prevent the introduction of extraneous de-

Jan 1, 1964

By M. D. Arnold, P. B. Crawford, P. C. Hall

An experimental study has been made to determine the optimum flooding pressures for four different oils. The oil formation volume factors ranged from 1.08 to 2.13, and solution gas-oil ratios ranged from about 200 cu ft/bbl to 2,250 cu ft/bbl. Viscosities ranged from 0.38 to 0.95 cp at the respective bubble points of the fluids and from 0.7 to 20 cp at atmospheric pressure. Water floods were conducted at various pressure levels from run to run. The recovery as a function of flooding pressure was found to be different for each fluid, with optimum gas saturations ranging from 7 up to 35 per cent. The data indicate that substantially higher recoveries may be obtained if water floods are conducted at an optimum pressure and that this optimum pressure is a function of fluid properties. The same core was used for all tests, and the reproduction of saturations for various runs indicates that wettability in the predominantly water-wet core did not change. INTRODUCTION A paper was presented by Bass and Crawford' which described an experimental study of the effects of flooding pressure and rate on oil recovery by water flooding. This work was conducted using high-pressure models operated in a manner similar to that of an actual reservoir, with gas saturations being obtained by a solution-gas-drive mechanism. They found that the greatest oil recovery was obtained for the system studied by flooding in the presence of a 5 to 7 per cent gas saturation. Another experimental study simulating field conditions was presented by Richardson and Perkins.' They used an unconsolidated sand pack containing kerosene-natural gas fluid and interstitial water. They flooded at various pressures and flooding rates. For their system it was found that the recovery was independent of the pressure level at which the water flood was performed. Kyte, et al," found that oil recovery by water flooding was increased as the free gas saturation at waterflood initiation was increased. However, after the initial gas saturation was increased above 15 per cent, the increase in oil recovery tended to level off. All of their runs were made at the same pressure using a light oil saturated with helium. The desired gas saturation was obtained by injecting helium into the core. Dyes' made calculations which showed that an optimum gas saturation of 12 to 14 per cent may result in an increase in oil recovery of 7 to 12 per cent over that obtained by flooding at the bubble-point pressure. Others have also found that the presence of a free gas saturation may increase the waterflood oil recovery. In each case shrinkage was small and changes in fluid properties with respect to pressure were small. A careful review of the literature reveals that at the present time there is a wide difference of opinion on the factors affecting waterflood recoveries. This diversity of opinion is probably due to the fact that very little research has been done which has taken into account the many variables existing in an actual field being water flooded. Since the literature contains little information on high-pressure waterflooding studies using various types of reservoir fluids, it was believed appropriate that such a study should be made. EQUIPMENT AND PROCEDURE All tests were made using the same consolidated sandstone core. Torpedo sandstone was used to turn a cylindrical core 13.5-in. long and with a 2.92-in. average diameter. The core had a porosity of 28 per cent and a permeability to brine of 146 md. This brine was made up by adding 20,000-ppm sodium chloride and 30,000-ppm sodium nitrite to distilled water. This was used as connate water and flooding water. No fresh water was ever brought in contact with the core, as tests showed the sandstone contained argillaceous material which swelled in the presence of fresh water and plugged the stone. The core was sealed in a section of 6-in. N-80 tubing with Woods metal filling the annulus. The core was mounted horizontally; an injection well was placed in the center of one end and a production well in the center of the other. Pressure control was maintained by placing a back-pressure regulator (upstream control) on the producing well. The "live" oil was stored in a separate bottle and water was injected into this bottle to displace the oil for saturating the core using a two-cylinder standard-proportioning pump. This same pump was used for water flooding the core at a constant rate. This system was enclosed in water jackets and the temperature was automatically main-

By S. M. Purdy

Under certain conditions of hot rolling and air cooling from the hot-rolling temperature, bars of a high carbon (0.40 pct C) chrome-nickel austen-itic alloy were found to show magnetism even though no ferrite or martensite could be detected by microscopic or X-yay methods. The appearance of magnetism in such alloys may come from chromium impoverishment of the austenite grains near the precipitated carbide particles. SPORADICALLY, hot-rolled bars of Silchrome 10, an exhaust valve steel, have been found to be magnetic. Because of the analysis of the alloy—0.40 pct C, 18 pct Cr, 8 pct Ni, 3 pct Si —magnetism is unexpected. Preliminary investigation showed neither martensite nor ferrite to be present; only austenite and Cr23C6. Since a literature search was fruitless, a brief study was made of the appearance of magnetism in this alloy. The only basic difference between the two heats is the nitrogen content. Permeability was measured using a Severn magnetic gauge. This instrument consists of a magnet mounted on a counterbalanced arm. A set of calibrated plugs is placed in contact with one pole of the magnet. The specimen is placed close to the other pole of the magnet. If the specimen pulls the magnet away from the plug, it has a permeability greater than that marked on the plug. This technique is swift and reproducible. Previous experience has shown that the permeabilities obtained corresponded to those obtained on a permeater with a field strength of 100 oe. Specimens from both heats were annealed at temperatures between 1700 and 2300°F. One set of specimens was water cooled and another furnace cooled. All the water-quenched specimens were non-magnetic; the furnace cooled ones were magnetic as shown in Table I with no difference being observed between the two heats. Microstructural examination of the specimens showed the expected increase in carbon solubility with increasing temperature. Carbide solution was complete at 2200°F. The specimens heated to 1900°F or below showed some carbide precipitation from the hot-rolled structure. A furnace cooled specimen from a given temperature showed less carbide out of solution than the water-quenched specimen from the next temperature below; e.g., the specimen furnace cooled from 2100°F showed less carbide out of solution than the water-quenched specimen from 2000" F. These studies indicated that the appearance of magnetism was not related to the quantity of carbon in or out of solution and it was related to precipitation at temperatures below 1700" F. A set of samples annealed and water-quenched from 2100° F was aged for 4 hr at temperatures between 1000" and 1600°F; all were non-magnetic. A second set of samples, similarly annealed, was aged 1 to 24 hr at 1200°F with the results shown in Table II. None of the latter set of specimens showed magnetism until they had been aged about 8 hr. Magnetism was quite strong after aging 24 hr. X-ray diffraction studies on several of the magnetic specimens showed that the austenite had a lattice parameter of 3.58A and that the carbide was Cr23C6. Several of these samples were electrolytically digested in 10 pct HCl in ethanol, with a current density of 0.1 amp per sq cm. None of the particles in the residue were magnetic. Accidentally, one cell was run at 1 amp per sq cm; e.g., magnetic particles were found in this residue. After careful separation, the magnetic particles were mounted on a quartz fiber and their diffraction pattern determined using a 5.73-in. Debye-Sherrer camera with CrK radiation. These particles showed a fcc structure with a lattice parameter of 3.57A. Prolonged exposure, up to 16 hr, produced no other lines on the film. The following facts seemed to be established at this time: 1) Austenite was the magnetic phase. 2) Neither ferrite nor martensite could be detected. 3) Magnetization could be produced by aging at 1200°F. One explanation of these data is that the carbide precipitation impoverishes the region immediately around the carbide particle of carbon and chromium and increases the proportion of nickel. All of these serve to increase the Curie temperature of the region around the carbide particle. If the composition change is enough, the Curie temperature will rise above room temperature. If the volume of the affected region is great enough, the magnetism will become detectable. At low aging temperatures, composition changes are great enough but the overall volume of impoverishment is quite small

Jan 1, 1962

By W. R. Hudspeth

AS an outgrowth of its spodumene recovery operation at Kings Mountain, N. C., Foote Mineral Co. has been recovering a heavy mineral by-product. Foote leased this idle plant in 1951, reactivated it, using a new spodumene recovery process, and purchased plant and properties in October 1951. While the operation at Kings Mountain is primarily concerned with the production of spodumene concentrate, pilot plant work determined that the pegmatites also contained heavy minerals including cassiterite. Plans were made to recover the heavy minerals as a by-product and the flowsheet incorporated these facilities when the mill was modified for the new spodumene recovery process. The orebodies consist of spodumene, feldspar, quartz and mica. Apatite, tourmaline, and beryl are present in small quantities. The wall rock is pre- dominantly hornblende shist. The heavy minerals, including cassiterite, columbite, pyrrhotite, monazite, pyrite, and rutile represent about 0.2 pct of the ore. The fine-grained heavy minerals are disseminated throughout the dikes, apparently unassociated with the spodumene. The pegmatites are quarried and secondary breakage is by mud-capping and block-holing. Power shovels load into trucks transporting the ore to a coarse ore bin. A Telesmith 10x36-in. apron feeder delivers the ore to an 18x36 in. Traylor Jaw crusher adjusted to discharge -3 in. product to a primary conveyor. The conveyor delivers to a 4x5-ft Tyrock single deck vibrating screen using 3/4 in. cloth. The screen undersize is elevated to the crushed ore bin. Screen oversize goes to an Allis-Chalmers Hydrocone Crusher fitted with 4 in. concave and set to deliver approximately 66 pct minus 3/4 in. The crusher discharge returns to the primary conveyor. The crushing and screening installation has a capacity of about 60 tons per hour. Spirals The crushed ore is delivered at a rate of 350 tons per day to two 6x8-ft Hardinge Pebble Mills, equipped with 20 mesh Ton-Cap trommel screens. The screen oversize is pumped to a 12-in. hydroclone for primary desliming. The hydroclone underfl spirals. There is no heavy mineral loss in the hydro-clone overflow. The spirals bank consists of eight 5-turn Model 24-A Humphreys Spirals. The top port and the last four ports of each spiral are blanked, the remaining nine port splitters are adjusted to remove about 5 pct feed weight. The heavy mineral rougher concentrates are upgraded on a Deister Overstrom table. The spiral concentrates contain approximately 4 pct heavy mineral, and the spiral reject, which goes to another section of the plant for spodumene recovery, contains about 0.03 pct heavy mineral. There is an interesting feature in the spirals installation. An adjustable splitter mounted on the discharge boxes splits out a mica fines product containing very little heavy mineral. The mica product is cleaned by spiralling and screening. Thus the spirals recover two products; mica, and a heavy mineral rougher concentrate. Table Treatment The rougher spiral concentrate goes to a Deister Plato table, modified to receive a Deister-Overstrom No. 6 rubber cover with sand riffles. The table is operated with a 5/8 in. stroke, 270 strokes per minute, and a slope of 1/2 in. per ft from feed to tailings side. There is no slope adjustment from motion to concentrate end. Wash water consumption is relatively high, since the large spodumene grains tend to report with the fine heavy minerals. A middling band about 4 in. wide is maintained in order to produce clean concentrate. The middling, representing about 10 pct of table feed, is recirculated by air-lift. A band of concentrate grade coarse spodumene occurs just below the middling. This is removed and delivered to concentrate storage. The table tailing, containing approximately 0.7 pct heavy minerals, is returned to the spodumene feed preparation circuit. The heavy mineral table concentrates are approximately 45 pct cassiterite, 33 pct columbite, 14 pct pyrrhotite and 8 pct monazite, together, with some rutile, pyrite, and copper from blasting wire. Concentrate is collected at 24 hour intervals. and dried. If the concentrate remains in wet storage appreciably longer surface oxidation takes place which seriously interferes with the subsequent magnetic separation process. About 150 lb of heavy mineral concentrate is produced per 24 hours and shipped to the company's plant at Exton, Pa. for final separation into tin and columbium concentrates.

Jan 1, 1952

By S. S. Cole, H. Sigurdson

The melting points of mixtures of titanium dioxide and other titanates have been reported to a limited extent as binary systems and some results have been reported in conjunction with silicon dioxide. The limited data indicated that a low melting zone might exist in a ternary or quarternary system of CaO-MgO-TiO2-Al2O3, since the eutectics reported in the binary systems of MgO.TiO2-TiO2, Al2O3TiO2-TiO2 and CaO.TiO2-TiO2 were of about equivalent composition of TiO2. The importance of a low melting region with a high titanium oxide value in the developing of high titanium slags is fully appreciated by metallurgists. A comprehensive study was undertaken to establish basic information on low melting titanate mixtures. Compounds in the System CaO-MgO-TiO2 -Al2O3 Preparation of the titanates of CaO, MgO and Al2O3 confirmed reported data that the following compounds could be formed under oxidizing conditions in solid state reactions: CaO.TiO2, 2MgO.TiO2, MgO.TiO2, MgO-2TiO2 and Al2O3.TiO2. Under oxidizing conditions it was not possible to react MgO with TiO2 in mol ratios higher than 1:2 without having unreacted TiO2 in the product, nor was it possible to form calcium or aluminum titanates with a higher mol ratio of CaO and A12O3 to TiO2 than 1:1. By fusing mixtures of Al2O3-TiO2 and MgO-2TiO2 together a series of solid solution products were obtained, which showed an MgO-2TiO2 X ray diffraction pattern shifted to smaller inter-planar spacings. Range of Investigation Data in the literature1 indicated that these titanates had melting points which ranged from 1645 to 1860°C. Since the purpose of the study was primarily to provide useful data for smelting titaniferous ores, the work was restricted to the zone of the system which would include the crystalline phases present in the slags. The explored limits were bounded by CaO.-TiO2-MgO.TiO2-TiO2in the base plane. The system was extended to a fourth component Al2O3-TiO2 since many titaniferous ores contain appreciable amounts of A12O3. In the quarternary system CaO-TiO2-MgOTiO2-TiO2-Al2O3TiO2, a tetrahedron was used to represent graphically the components, with one component at each point of the tetrahedron. In the base plane, mixtures were pre- pared to represent fairly uniform changes of composition over the desired range expressed on a mol percent basis. Al2O3-TiO2 was brought into the system in increments of 10 mol pct up to 40 pct. Mixtures were then made up for each of the Al2O3-TiO2 planes. The ranges employed were then represented by five planes cutting the tetrahedron at 10 mol pct Al2O3-TiO2 intervals. Choice of Equipment Although it was necessary for all smelting work to be done in a strongly reducing atmosphere, it was decided that all melting point determinations should be done under oxidizing conditions. This decision was prompted by the fact that TiO2 reduced to lower oxides in a reducing atmosphere with consequent changes in melting points. The use of a micropyrometer and a platinum strip furnace for work of this type is adequately described in the literature.2 With modifications, equipment was selected which provided a fairly rapid method of determining melting points of refractory oxides. The final assembly, which was used, consisted of a platinum strip furnace and a Leeds and Northrup disappearing filament optical pyrometer attached to a special telescope which magnified the sample about 20 diam. The furnace assembly consisted of a platinum strip 0.005 X 0.3 X 2.2 in. mounted on brass posts on a refractory base. This was enclosed in a black steel shell which had a 2-in. opening at the top for purposes of sighting the telescope. The platinum strip was in series with two 0.04 ohm nichrome resistors connected to the secondary of a 2 kw transformer which supplied 17 volts. The current to the primary was varied

Jan 1, 1950

By Sandford S. Cole, John S. Breitenstein

The recovery of over 80 pct of the vanadium values in titaniferous magnetite from Maclntyre Development,Tahawus, N. Y., was accomplished by an oxidizing roast with Na2O3-NaCI addition. Process description is given for leaching of roasted ore and precipitation of V2O5 and Cr2O8 from leach liquor. THE exploration and development of the Mac-Intyre orebody at Tahawus, N. Y., by the National Lead Co. provided a source of vanadium. Analyses of various composite sections of the drill cores of the MacIntyre orebody were made to establish whether or not the vanadium was constant throughout. Ten drill cores were sampled as 50 ft sections, crushed, and a portion magnetically concentrated. The head and concentrate were analyzed for total iron and vanadium. The results on the concentrates indicated that the vanadium is associated with the magnetite and maintains a close ratio to the iron content. The nominal ratio of 1:25:140 of V: TiO2:Fe was found to exist in the concentrates. Typical value for the vanadium in the magnetite both from laboratory concentration and mill production is 0.4 pct. The recovery of vanadium from the magnetite was investigated in 1942 to 1943. The research program encompassed both laboratory and pilot-plant work on sufficient scale to provide adequate data to establish the feasibility of a full scale plant. The recovery of vanadium from various ores has been reported in the literature and has been the subject of many patents. The literature dealing with recovery from titaniferous ore by roasting is quite limited. Roasting with alkaline sodium chloride, sodium chloride or alkaline earth chlorides, and sodium acid sulphate have been claimed in various processes as effective means.1-8 The reduction of the ore, followed by acid leaching, was another method proposed.'-' "he use of various pyrometallurgical processes for recovery of vanadium in the metal or in the slag has also been extensively investigated, but the results had little application to the problem."-" The separation of vanadium values from subsequent leach liquors and vanadium-bearing solution has been the subject of a considerable number of papers and patents. The most practical is by hydrolysis at a pH of 2 to 3 by acidifying a slightly alkaline solution. Data on solubility of V²O5 and V2O4 in water and in dilute sulphuric acid indicated a solubility of 10 g per liter in water.'" Laboratory Results Magnetite Analysis: Adequate stock of magnetite was provided so that the laboratory and pilot-plant operation was on ore representative of the mill production. The ore was analyzed chemically and examined by petrographic methods to ascertain whether the vanadium was present in combined state or as an interstitial component between grain boundaries. No evidence was obtained which would indicate that the vanadium was in a free state as coulsonite.15 The analysis of the ore was as follows: Fe²O³, 47.4 pct; FeO, 29.1; TiO,, 10.1; V, 0.40; and Cr, 0.2. The screen analysis of the ore on the as-received basis was: -20 +30 mesh, 28.8 pct; —30 +40, 18.9; -40 +50, 9.7; -50 +60, 15.1; -60 4-100, 5.9; -100 + 200, 11.2; -200 +325, 3.7; and -325, 7.2. Roasting Conditions: The prior practice indicated that a chloridizing roast with or without an alkaline salt had been effective on other titaniferous magnetites. On this basis roasts with additions of sodium chloride, sodium carbonate and mixtures thereof were investigated varying the roasting temperature between 800" and 1100°C. Since the ore had shown no segregation or concentration of vanadium, the influence of particle size on the freeing of vanadium by the reagents during roasting was determined. The initial work was on silica trays in an electric resistance furnace with occasional rabbling of the charge. Subsequently, the roasting was carried out in a small Herreshoff furnace to establish the influence of products of combustion on the recovery of the vanadium. The laboratory tests showed that this ore required an alkaline chloridizing roast, in conjunction with a reduction in particle size to less than 200 mesh. When roasted in air at 900 °C with 5 pct NaCl and 10 pct Na2CO³, over 80 pct recovery of the vanadium was attained as a water-soluble salt. The presence of alkaline earth elements gave detrimental effects and care had to be exercised to avoid any contamination of the ore or roast product by such materials. The solubilization of vanadium under the various conditions is given in a series of curves in Figs. 1 to

Jan 1, 1952

By T. A. Rickard

The dictionary defines 'jargon' as "barbarous or debased language". This description does not suffice. Quiller-Couch has said, it is "a kind of writing which, from a superficial likeness, commonly passes for prose in these days, and by lazy folk is commonly written for prose, yet actually is not prose at all ". The two main vices of jargon, he says, are "that it uses circumlocution rather than short straight speech", like the Babu who reported his mother's death by saying: "Regret to inform you, the hand that rocked the cradle has kicked the bucket ". Its other chief' vice is that "it habitually chooses vague woolly abstract nouns rather than concrete ones", like the newspaper statement; "He was conveyed to his place of residence in an intoxicated condition", instead of saying, "He was carried home drunk ". Jargon is "an infirmity of speech", it is not journalese, but akin to it. "Like respect- ability in Chicago, jargo4 stalks unchecked in our midst" and renders much technical writing ridiculous. It deals in periphrasis instead of going straight to the point, it loves the abstract rather than the concrete, it dabbles in words of sound rather than of meaning. Avoid it, despise it., if you purpose earnestly to write well. "III literature as in life he makes himself felt who not only calls 4 spade a spade but has the pluck to double spades and re-double." * Jargon is rampant in technical publications. Catalogues and other advertisements are conspicuous offenders, but with these the critic is not concerned, except in so far as such 'write-

Jan 1, 1931

By J. L. Huitt, D. K. Lowe

This laboratory flow study covers propping agent transporl in horizontal fractures as influenced by the characteristics of the propping particles, fluid and fracture. Correlations are presented for the transport of particles, deposition and fingering through dunes. For particles moving individually, the article velocity in some cases may be low compared to the hulk average fluid velocity; in other cases, the particle velocity may exceed the bulk average fluid velocity. At high propping agent concentration or low fluid velocity, the likelihood of particle deposition and dune formation increases. Fingering and channeling accompany dune formation. Dune formation can be limited to the outer region of tile fracture by adjustment of pump rate, fluid viscosity and particle concentration. A method is presented by which pump rate and fluid vircosity may he selected to control dune formation for given propping agent concentrations. INTRODUCTION The area and flow capacity of a fracture predominantly control the stimulation derived from hydraulic fracturing. Generation of fracture area exposes the rock matrix to a flow channel; the flow capacity of the fracture controls the flow through the channel to the wellbore under a given pressure gradient. This paper pertains to an important factor that affects the flow capacity of a fracture: the transport of the propping particles in the fracturing during the hydraulic fracturing treatment. With a better understanding of particle transport, fracturing treatments may be designed to obtain more effective distribution of the propping agent. RELATED STUDIES In recent years, the approach to obtaining high fracture flow capacity has been partial monolayer propping. Three papers1-3 have dealt with various facets of this concept. However, in some papers dealing specifically with particle transport, the results generally indicate that packing the propping particles within a fracture is inevitable. In early work&apos; with a sand slurry flowing in a pie-shaped flow cell, qualitative results showed that particle deposition in a fracture during transport should be expected. In other studies"" correlations based on slurry flow in vertical linear flow cells showed particle deposition was primarily a function of fluid velocity. In a recent paper the rate of sand advance (movement of dunes) in a horizontal linear flow cell was discussed. Subsequently, a method was given for designing fracturing treatments, based on the extent of sand advance during a treatment. In this work, particles whose diameter was small relative to the fracture width (i.e., dl W< 0.2) were primarily considered. The particle transport approaches that of sand transport in a stream bed. The industry trend, however, has been to use larger size propping agents. That is, particles with diameters approaching fracture widths predicted by Perkins and Kern.&apos;&apos; This trend has led to generally better fracturing treatments with particle relative sizes in the range 0.2 <d/W < 1.0. In all of the particle transport studies previously cited, the fluid characteristics were discussed. However, little or no attention was given to the parzicle characteristics. In this study the particles. as well as the fluid and fracture, are considered. VARIABLES AFFECTING PARTICLE TRANSPORT In the exploratory work to set LIP a program for this study, it was noted that particles within a batch moved at different velocities in a horizontal linear flow cell. In some cases the particles moved independently. In others there was particle interference. Also, particle deposition and fluid fingering occurred in some cases (Fig. 1). As a result of observing the easily distinguishab&apos;le types of particle movement, the particle variables as well as the fluid variables were included. In addition, the fracture width was included because the fluid velocity profile surely must affect the movement of the larger particles differently than that of smaller particles. Particle variables were density, diameter (based on a sphere of equal volumetric displacement), long and short dimensions, drag coefficient and velocity. Fluid variables were density, viscosity, yield shear strength and velocity. Fracture variables were the width and the angle between the plane of flow and a horizontal plane. Two factors which may prove to be important were omitted because of the complexity resulting from the parameters cited above. These are the fracture-surface roughness and gross irregularities that cause changes in the direction of flow. EXPERIMENTAL EQUIPMENT AND PROCEDURE A linear parallel-plate flow cell (Fig. 2) was used in the study. Lucite plates 1 in. thick formed the test section and permitted a viewing section 11 ft long and 10 in. wide. Spacing between the plates was adjusted with shims to give fracture widths 1/36 to 1/2 in. The model could be tilted to an angle of 45" with the horizonltal plane. Fluid flow was maintained with a Moyno positive displacement pump. Single particles were placed in the flow stream with a particle injector. A .4 - in. diameter plunger moving at a constant rate in a

Jan 1, 1967

By G. Y. Chin, W. L. Mammel

The problem of selection of the active slip systems for a crystal undergoing an arbitrary strain has been analyzed by Taylor and by Bishop and Hill. The Taylor analysis is based on a principle of&apos; virtual work, and involves finding, among numerous cotnbinalions of slip systems that satisfy the imposed strain, the combination in which the sum of the glide shears is a minimum. Previously, Taylor has treated the case of axisymmetric flow when slip occurs on (111)(110) (or {110)(111)) systems only. His analysis has now been extended by computer methods to the cases of slip on {112}(111) and {123)(111) systerns and of mixed slip on {110), {1 12), and (123) planes with a common (111) slip direction, all of which are important in the deformation of bcc crystals. The results are computer-plotted as contours of the ratio of the floe strength to the critical resolved slzea-r stress for slip, for axial orientations distributed throughout the standard stereographic triangle. Implications of the computer results to texture develop,merit, texture hardening, and dislocation theories oj work hardening are discussed. WhEN a single crystal is extended in the usual tension test, the lateral dimensions can change relatively freely. In this case, the glide shear produced by slip on a single slip system is sufficient to accommodate the (tensile) deformation. Since slip is governed by a critical resolved shear stress law (the Schmid law&apos;), the single active slip system is one for which the stress, resolved on the slip plane and in the slip direction, is the highest among the several equivalent slip systems. This amounts to saying that a value M = U/t = y/~ is a minimum among the equivalent systems, where M is the inverse of the familiar Schmid factor (a and E refer to tensile stress and strain, and T and y refer to resolved shear stress and shear strain). A grain embedded in a polycrystalline aggregate, on the other hand, cannot freely change its shape due to constraint from its neighbors. In this case, slip from five independent slip systems (to accommodate five independent strains) is generally required.&apos; Based on the principle of virtual work and assuming that the critical resolved shear stress for slip is the same for all systems, Taylor hypothesized that, among all combinations of (five) slip systems which are capable of accommodating the imposed strain, the active combination is that one for which the sum of the absolute values of the glide shears is a minimum. Again, this is equivalent to saying that the value of M = CjlyjI/ is a minimum, in analogy to the single slip case. Taylor aminimum,analyzed the case of {111)(110) slip for fcc metals, and applied the analysis to crystals undergoing axisymmetric flow, that is, the same macroscopic shape change as the poly crystalline aggregate under uniaxial tension (or compression). For the twelve equivalent {111}( 110) slip systems, there are 384 independent combinations of selecting five systems to satisfy the five independent linear equations of imposed strain.4 Taylor calculated the value of M for each combination* and obtained the active com- *A number of the independent combinations were omitted from consideration in Taylor&apos;s original work (see Ref. 5). bination (minimum M) for a number of axial orientations distributed throughout the standard stereographic triangle. Later work by Bishop and Hill*&apos;8 showed that Taylor&apos;s least-shear hypothesis was equivalent to a maximum work principle which they advanced. Using the simplified Bishop and Hill method for {111)(110) slip, Hosford and Backofen&apos; obtained detailed contours of constant minimum M for the same axisymmetric flow case. In contrast to {111}(110) slip in fcc metals, slip in bcc metals is generally described as occurring on {ll~)(lll), {112)(111), {123) (111) systems as well as mixed slip composed of all three. Since the direction cosines of the slip plane normal and the slip direction enter as a product in the Taylor analysis, the Taylor solutions of ,M for {110)(111) slip are identical to those for { 111} 110) slip. The other three cases of slip, however, have not been solved. In view of the numerous combinations of slip systems involved in the calculations, the Taylor analysis is clearly oriented toward computerized solutions. THE TAYLOR ANALYSIS In order to obtain the active combination of (five) slip systems by solving for the minimum value of M = we first express the (small) strain components E,, with respect to the cubic axes 1, 2, 3 ([loo], [010], [001], respectively) of the crystal, in terms of the sum of the glide shears yj from slip systems j: where n, and n,j refer to direction cosines of the slip plane normal, and dri and dsi to direction cosines of the slip direction, of slip system j, all referred to the cubic axes. In practice, the strain components are given with respect to the specimen axes X, y, 2. These components are readily converted to ers through the tensor transformation where irk and 1,~ are the direction cosines between the two sets of coordinate axes.

Jan 1, 1968

By A. M. Rowe, I. H. Silberberg

A computer program was written to predict the phase behavior generated by the enriched-gas-drive process. This program is based, in part, on a new concept of convergence pressme, which is then used to select vapor - liquid equilibrium ratios (K-factors) for performing a series of flash calculations. The results of these calculations are the equilibrium vapor and liquid phase compositions which define the phase envelopes. The program was used to predict phase envelopes for 11 different hydrocarbon systems on which published experimental data were available. The predicted and experimental results compare favorably. INTRODUCTION The reservoir engineer is frequently faced with the problem of predicting what will happen if gas is injected into a reservoir. One aspect of this, general problem is predicting the phase changes that will occur when a non-equilibrium gas displaces a reservoir fluid. In particular, a "dry" gas, upon displacing a volatile oil, will pick up intermediate components from the oil. On the other hand, a "wet" gas, containing a high concentration of intermediates, will lose some of these components to a relatively low-gravity, non-eouilibrium1- crude. It is this latter Drocess. occurring in the enriched-gas displacement, which is treated in this paper. In the past, these phase changes have been determined experimentally and the results incorporated into various modifications of the Buckley-Leverett analysis.112 Such experimental work is time consuming, and the results are sensitive to numerous experimental errors. With the wide availability of high-speed digital computing equipment and numerous correlations pertaining to the vapor-liquid equilibria of hydrocarbon systems, it is now practical to calculate such phase behavior. This paper describes a computer program for performing these calculations. THE ENRICHED GAS DISPLACEMENT PROCESS Experimental results have shown that oil recovery can be significantly increased by enriching the displacing gas with intermediate hydrocarbon c0m~onents.3 The essential features of the phase behavior generated by this enriched-gas-drive process are commonly illustrated with ternary diagrams such as Fig. 1.4 In this figure, Gas D, which contains a high concentration of intermediate hydrocarbons with respect to the undersaturated Crude A, is injected into the reservoir. When D contacts A, gas goes into solution until the oil becomes saturated (Point. B). Further contacting of Gas D and saturated Oil B results in a Mixture C which separates into Vapor Y(c) and Liquid X(c). Liquid X(c) is contacted by additional Gas O, resulting in Mixture E which separates into Vapor Y(e) and Liquid X(e). Repeated contacts of the liquid by the injected gas will eventually result in Liquid X(4 of maximum enrichment existing in equilibrium with Gas Y(d). The equilibrium tie-line X(4 Y(4, when extended, passes through the Point D representing the enriched injection gas. For systems of more than three components, the predicted equilibrium states are dependent upon not only reservoir temperature and pressure, but also the compositions of the crude oil and injected gas. If the gas is sufficiently enriched, a miscible displacement is generated. Line If is tangent to the phas,e envelope at the critical point (Point Z) and represents the limiting slope of the tie-lines as the critical state is approached. Point 1 therefore represents the minimum enrichment of injection gas required to generate a miscible displacement. Point G represents the minimum enrichment required for initial miscibility of the injection gas with Crude A. Accra has presented a method to be used for prediction of oil recovery by the enriched gas displacement process.l To develop the phase behavior data needed, he designed the experimental procedure described in the following quotation from his paper: The original liquid was contacted by a volume of displacing gas and allowed to come to equi-

Jan 1, 1966

BENEFICIATION AND CONCEN&apos;TRATION; FROTH FLOTATION U.S. 4,069,144 - In the froth flotation beneficiation of phosphate rock using at least an acid flotation step and an amine flotation step, the crude ore is dried either during transport or storage or by air circulation immediately prior to classification, and the dried ore mixed with water and classified prior to flotation The preliminary drying substantially eliminates the usual problems with clay slimes G. C. Johnson and A. E. Schweizer, assigned to Mobil Oil Corp. Jan 17. 1978. 4 pp (209- 11) . Filed 1 1-19-75 (633,535).

Jan 1, 1979

By E. Parthe, J. P. Pemsler

TWO references can be found in the literature concerning the ternary system uranium-tantalum -carbon. C. H. Schramm, P. Gordon, and A. R. Kaufmam reported in this Journal1 on the existence of the ternary carbide "UTa10C4." They also report a second face-centered-cubic carbide phase in the system U-Ta-C, which they denote as "white" phase. The publication includes a reproduction of a Debye-Scherrer powder photograph of "UTa10C4" and a list of the sin2 0 values of this compound. H. Nowotny, R. Kieffer, F. Benesovsky, and E. Laube2 report the complete solid solubility between TaC and UC with a B1 structure. Since the ternary phase "UTa10C4" lies very close to the binary system Ta-C, and the established phase Ta2C with a C27 structure had not been mentioned by Schramm et al., it was thought worthwhile to compare the patterns of "UTa10C4" with Ta2C. The strong lines of the given pattern were indexed on a hexagonal unit cell with the approximate cell parameters a = 3.09A and c = 4.93A. These are the values reported for Ta2C in the literature.3 The correspondence between reported and calculated sin2 8 values may be studied in columns 4 and 5 of Table I. The agreement is fair, but there remain weak lines which cannot be assigned to the hexagonal unit cell. The extinctions correspond to those of Ta2C, but a comparison of the calculated and observed intensities of the Ta2C lines (columns 2 and 3) with those of the corresponding lines of "UTa10C4" (column 6) show occasional deviations. To clarify this situation, four new samples were prepared. The compositions of these samples were chosen in such a way as to observe any existing ternary phase in the system TaC-UC-U-Ta. The samples were arc melted in an inert atmosphere and annealed at 1100°C for 36 to 72 hr. Compositions of the samples determined by chemical analysis are given along with the results of X-ray investigations in Table 11. The following conclusions may be drawn from these data: a) No ternary compounds occur in the section TaC-UC-U-Ta of the ternary system Ta-U-C. The "white" phase reported by Schramm et al. corresponds to a solid solution between TaC and UC. Only one three-phase field exists: C27-B1-A2. (Both tantalum and ? uranium have body-centered-cubic A2 structures.) The a uranium observed in sample 4 transformed from the high-temperature ? phase on cooling. b) The phase Ta2C with a C27 structure extends into the ternary system. "UTa10C4" is, therefore, a uranium-enriched Ta2C. Good agreement is ob-

Jan 1, 1960

By Arthur Baker

The orebodies are long bedding-plane lenses of chalcopyrite and sphalerite, associated with garnetite masses. Most of the orebodies are within a 50-ft thickness of Cambrian limestone; other Paleozoic limestones and dolomites are locally metamorphosed but only slightly mineralized. Pre-mineral faults are numerous, but shallow folds were the main ore-localizing structures.

Jan 12, 1953

By B. N. Whittaker, C. J. Bonsall

The paper gives an account of the factors influencing the stability of coal mining tunnels and goes on to examine various bases of support and lining design of such tunnels in relation to geological and mining conditions. The different types of coal mining tunnels are compared and comments made regarding the factors influencing their choice and basic design features. Foremost amongst design considerations is the design method and special thought has been given to the validity of different formulae as related particularly to UK Carboniferous geological conditions. Emphasis is placed on the role of the stress field in relation to tunnel stability. Shape of mining tunnel has been investigated and conclusions drawn regarding the merits of different commonly excavated geometries from a stability point of view. Detailed design aspects pertaining to determination of tunnel lining strength and resulting stability are critically examined and current design methods are compared. Data from UK mining tunnels are included as the basis for comparison of these design methods.

Jan 1, 1983

By Dan White

Stress measurements taken at Climax, El Teniente, Henderson, Lakeshore and Urad mines indicate that the deposits cited are generally subject to an inclined stress field that can be highly anisotropic. The horizontal stresses vary significantly between deposits and exceed the measured vertical stresses at Henderson and El Teniente mines. Experience at Henderson, Lakeshore and Urad mines is used to suggest a relationship between cavability and the premining stress. In poorly fractured, high strength rock masses, high stresses can assist the caving process. In well fractured, low strength rock masses, caving is ensured by breakage along fracture planes. The relationship between ground weight and the premining stress is also investigated. Ground weight problems appear to be most related to the competence of the ground rather than the magnitude of the premining stress as supported by experience at Climax and El Teniente mines. Recent developments in block caving favor the extraction of poorly fractured, high stressed rock. Accompanying the development of these deposits are new problems such as rock bursts. These pose particular problems to mine designers in that no experience currently exists in combatting rock burst problems at block caving mines.

Jan 1, 1984

By W. J. Keep

Cast-iron, or pig-iron, is iron which contains all the carbon that it could absorb during its reduction in the blast-furnace. As is well-known to chemists, carbon exists in cast-iron in two distinct forms. In chemical union, as "combined" carbon, it cannot be discerned, except as it may increase the whiteness of the fracture, in so-called white iron, or carburetted iron (sp. gr. about 7.72). Carbon mechanically mixed with the iron as graphite is visible, varying in color from gray to black, while the fracture of the iron, according to the color and amount of graphite it contains, ranges from a light to a very dark gray. The specific gravity of graphitic iron is about 7.40. Silicon in cast-iron is evidently to some extent combined with the iron and the carbon ; but whether it exists also in a form corresponding with graphitic carbon, mechanically mixed with the remaining mass, is a question still in dispute, which we have not in this paper endeavored to answer. It has been the opinion of many writers on metallurgy, and many foundry men, that silicon is in every way injurious to cast-iron. Yet

Jan 1, 1889

By W. G. Oldham

By monitoring the capacitance of abrupt p-n junctions it is possible to follow the motion of substitu-tional impurities. A p-n junction is formed by growth of silicon from an Al-Si alloy on an n-type silicon sutstrate at a temperature in the range 650" to 750°C. The diodes are etched and sorted for consistent V-I and capacitance behavior and separated into groups. After irradiation, each set with an unirradiated control set is annealed to various temperatltres for 10 mm and the capacitance remeasnred. For very heavily doped diodes (-1019 per cm3) observable capacitance changes occur at neutron fluences of >1015 per cm 3. The results may be interpreted in seceral ways, the simplest of which is to assume simple vacancy diffiision with small impurity -racancy interaction (at the diffiision, i.e. annealing temperature of 700°C). In this interpretation it is found that a typical vacancy makes about 1024 jumps and travels about 10-5 cm before disappearing. The required effective annihilation center density is about 1016per cm3 . Estimates of the vacancy-impurity interaction can be made and the above numbers correctecl, but for arty reasonable assumptions the number of vacancy sinks connot be much reduced. ThE complete annealing of radiation induced defects requires that the vacancies and interstitials either recombine, travel to the surface of the crystal, or travel to an annihilation center, e.g., a dislocation or a precipitate. If the interstitial is ignored except as a possible vacancy annihilation center at low temperatures,* the motion of vacancies may be followed by monitoring the corresponding motion of substitutional impurities. pfister2 has measured the enhanced diffusion of impurities in an asymmetric shallow diffused silicon structure by optically monitoring the junction movement. In these experiments we monitor instead the capacitance of a heavily-doped deep symmetric p-n junction. The sensitivity for this method is several orders of magnitude higher than in Pfister&apos;s technique, mainly for two reasons: 1) the junction capacitance of a heavily-doped junction is sensitive to impurity movements in the Angstrom range* com- and 2) the junction may be far from the surface which is a major annihilation center for vacancies. THE EXPERIMENT A p-n junction is formed by the epitaxial growth of silicon from an A1-Si alloy on a silicon substrate at a temperature in the range 650° to 750°C. The technique used is the transport of an aluminum melt through a temperature gradient.4 A 10-mil thick aluminum disc is sandwiched between two silicon wafers and the assembly raised to about 650° C with a temperature gradient normal to the sandwich. After about 8 hr the upper (hotter) wafer has been entirely transported and deposited epitaxially on the substrate wafer. The substrate and overgrowth resistivity are measured with a 4-point probe. The as-grown structure is lapped until the wafer is about 1/2 mm thick with the junction in the center. Contacts consisting of electroless nickel covered by electroplated rhodium are applied to both the p and n sides. The wafer is diced into square dice 1/2 to 1 mm on a side and the dice are etched sufficiently to remove approximately 1 mil of silicon. The following results are all for aluminum (p-side) and antimony (n-side) doped specimens with NA =3x 1018 per cm3, ND = 2-8 x 101A per cm3. The diodes are sorted for consistent V-I and capacitance behavior and separated into groups. After irradiation, each set with an unirradiated control set is annealed and the capacitance remeasured. A special capacitance measurement method is used to insure a low sensitivity to shunt resistance. The diode is voltage driven and the current monitored by a phase lock amplifier. It is possible to adjust the detector phase to a point where shunt resistances as low as 100 O have no effect on the reading. The diodes in this study have a capacitance in the range 1000 to 5000 picofarads and a shunt resistance in the range 100 to l06 O depending on the stage of the anneal. THEORY The theoretical problem is twofold: 1) The computation of the amount of impurity diffusion and hence the impurity profile resulting from a given amount of vacancy motion, and 2) the computation of the capacitance from the impurity profile. A comparison of the theoretical and experimental capacitance (in particular the changes) yields the amount of vacancy motion. Vacancy Motion. A convenient parameter to measure the total amount of vacancy motion is Jv the total number of vacancy jumps per unit volume. Although this parameter varies with position in the crystal, owing to vacancy annihilation at the surfaces, it is approximately independent of position far from the surfaces, i.e. in the junction region. In terms of the number of vacancies per unit volume Nv and the vacancy diffusivity given by

Jan 1, 1970

By C. S. Finney, John Mitchell

On December 31, 1959, there existed in the United States 15,993 slot-type coke ovens capable of producing 81,447,700 net tons of coke. These ovens were concentrated in 74 coke plants in 21 different states. As of the same date, there were 7448 beehive ovens in existence at 45 plants in the states of Pennsylvania, Virginia, West Virginia, and Kentucky. Total annual capacity of the existing beehive ovens was 4,368,800 net tons, but only 5148 ovens with a capacity of 3,131,600 tons were in operating condition. It is interesting to compare the average dimensions of slot-type ovens built during recent years with the 30 ft x 5 ½ ft x 16 ½ in. ovens erected at Syracuse, N. Y. in 1892. A composite oven built according to the average dimensions of all those erected between 1954 and 1958, for instance, would be 39 ft long. 12 ft high, and 18 in. in width. The coal capacity would be 16 tons as against the 4.4 tons which could be charged to the Syracuse ovens. Of the 15.993 slot- type ovens in existence at the end of 1959, by far the greater number were built by the Koppers Co. whose total of 11,280 ovens included 7891 Koppers- Becker and 3389 Koppers ovens. Of the remainder, there were 3260 Wilputte, 1350 Semet-Solvay, 63 Otto, and 40 Simon Carves ovens. By-product coke oven plants are usually classified either as furnace or merchant plants. According to the definitions used by the US Bureau of Mines, the former are "those that are owned by or financially affiliated with iron and steel companies whose main business is producing coke for use in their own blast furnaces. All other coke plants are classified as merchant. They include those that manufacture metallurgical, industrial, and residential heating grades of coke for sale on the open market; coke plants associated with chemical companies or gas utilities; and those affiliated with local iron works, where only a small part (less than 50 pct of their output) is used in affiliated blast furnaces." The annual coke capacity of the merchant plants during 1959 was 10,393,000 tons. However, the by-product oven of today is essentially an appurtenance of the iron and steel industry, rather more than 87 pct of total by- product coking capacity being concentrated at furnace plants. This was not always so. There was a time when the merchant plants played a much greater part in meeting the US demand for coke and gas. High noon for the merchant plants was reached during the early 1930&apos;s. By 1932 there were as many by- product oven installations being operated by the merchant sector of the industry as by the coke divisions of the iron and steel industry (44 of each), and in the same year the merchant plants produced 46.5 pct of all by-product coke made in the country. Since that time their contribution has drastically declined. In 1940 merchant plants were responsible for only 23.2 pct of total US production, and by 1950 their number had decreased to 30 plants which turned out 18.5 pct of the total by-product coke made. At the end of 1959 only 20 of the 74 existing by-product oven installations were merchant plants. They ac- counted for 12.5 pct of the year&apos;s production, or 6,849,786 net tons. This percentage has remained fairly constant since 1954. There are several reasons for the decline of the merchant coking industry. For example. On the grounds of economy, quality control, continuity of supply, and so on, the iron and steel industry usually prefers to control its own mines and carbonize its own coal at or near to the blast furnace rather than rely on independent operators for metallurgical coke. As the steel companies have enlarged their own coking facilities, so has the need for coke obtained from other sources declined. Furthermore, not only has the steel industry increased in self-sufficiency by building mare coke ovens during recent years, but it has also progressively improved the fuel efficiency of its blast furnaces. During the years 1947-49 the average coke consumption per ton of pig iron was 1892.8 lb. During 1958 the corresponding figure was 1613.4 lb. There are many individual furnaces where still better results are being obtained, and further reductions in the average may be expected. Perhaps the greatest threat to the merchant coking plant has been the fantastic increase in the use of natural gas and petroleum products for purposes which manufactured gas once served. So deadly has the com- petition from natural gas and oil been that it has almost eliminated by-product oven installations owned by public utilities. In the peak years of the early 1930&apos;s there were 23 such public utility plants. In 1960 only two were left. One of these, owned by the Citizens Gas and Coke Utility, was at Indianapolis, Ind.; the other was the plant operated by the Philadelphia Electric Co. at Chester, Pa. The non-utility merchant plants have also been sorely hit. With gas sales revenues reduced, domestic

Jan 1, 1961