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By J. Chipman, J. C. Fulton

Reduction of Si from slag to carbon-saturated iron is a very slow reaction. The rate is nearly independent of stirring but is accelerated markedly by increased temperature. In a slag containing 45 pct SiO2, 38 pct CaO, 17 pct Al2O3, the energy of activation is approximately 130 kcal. AT temperatures of 1450" to 1600°C occurring in the hearth of the iron blast furnace, chemical reactions are normally expected to proceed with great rapidity. Many of the reactions of iron making are indeed quite rapid and attain a condition corresponding roughly to a state of equilibrium before metal and slag are tapped from the furnace. The one conspicuous exception is the reaction by which carbon in the coke or metal reacts with silica in the slag to deliver silicon to the metal. From the practical viewpoint the slowness of this reaction is indeed fortunate, for if it were fast enough to reach equilibrium during the time of contact, it would be difficult to produce pig iron of the desired low silicon content. The one other reaction which under certain circumstances can also be very slow is the transfer of sulfur from blast-furnace metal to slag. In their laboratory exeriments on this reaction, Hatch and Chipman found that at least 5 hr were required for the attainment of equilibrium in a small crucible and that the reaction was not accelerated by increasing the rate of stirring. Subsequently it was shown2 that the presence of a reducible oxide MnO greatly inhibited sulfur removal, and finally that SiO2 itself3 under some conditions has sufficient oxidizing Dower to interfere with sulfur transfer. The kinetics of sulfur removal are now rather well understood,4 and it has been shown3,5 that when the silicon content of the metal is high enough to prevent reduction of SiO,, the desulfurizing process is rapid. Thus the observations of slow removal of sulfur are associated with the coincident slow reduction of SiO2, a process which transfers both silicon and oxygen to the metal phase. It is, of course, the oxygen which interferes with sulfur removal. An explanation of the slowness of the silica reduction reaction has not been obtained. The sulfur transfer process has been shown by the authors5 to be at least partially transport-controlled in the absence of interference by silica. Under the experimental conditions. 99 pct of the total transfer at 1600" was accomplished within an hour. Under similar conditions the reduction of silica may not proceed half-way to equilibrium6 in 8 hr. It is scarcely imaginable that silica reduction could be transport-controlled when the slag contains 44 pct SiO2 and the metal about 1.4 pct C. The experiments reported here show definitely that it is controlled by a slow step or steps in the chemical reaction. One might expect to find such a slow step in the breaking of SiO bonds or in the slow removal of oxygen by carbon at the slag-metal or metal-crucible interface or by bubble formation. EXPERIMENTAL PROCEDURE The experimental procedure used in this investigation was similar to that used in the study of the kinetics of desulfurization.5 A sketch of the graphite crucible and stirrer used in the current work is shown in Fig. 1. The quantity of metal and slag used was adjusted so that the slag-metal interface was located between the top and bottom blades on the stirrer; the top of the slag layer was above the top blade. The stirrer shaft was hollow down to the metal level and served as thermocouple well or optical pyrometer target. Stirring speeds of 0 to 500 rpm were used. A carbon monoxide atmosphere was maintained in the furnace by flushing the gas over the slag surface, except in run No. 220 when it was bubbled

Jan 1, 1960

By Orville R. Lyons

THE relative performance of coal washing equipment, or the effectiveness with which any type or make of equipment removes impurities from coal, has been most difficult to evaluate in the past. The most widely used yardstick is the Frazer and Yancey efficiency formula developed in 1922,1 but Yancey in a later article states that "washers treating coals of different density composition or operating at different densities of separation cannot be compared directly on the basis of this criterion."' Prior to and since 1922, a variety of other methods has been used for comparison purposes, including the distribution curve, the error area, and the "ecart probable" or probable error. Yancey and Geer in discussing these methods conclude, "Performance can be evaluated in a number of different ways, with the choice of the proper method to use being dictated by the objectives of the investigation and the data available."' It is true that performance can be evaluated in a variety of ways, but if the equipment is to be evaluated on an effectiveness basis, there should be only one universal comparison method. Varying methods have been used because one universal comparison method has not been found or developed. In the article previously quoted, Yancey and Geer state in clear terms the primary concept for a universal comparison method: "One of the simplest, and certainly one of the most obvious evaluations of washery performance is the quantity of sink material in the washed coal and the float material in the refuse. If the washery products are tested at the density at which the washing unit is operated, the sink in the washed coal and the float in the refuse represent material that has been misplaced." The quantity of misplaced material was used as a criterion of washery performance by Lincoln in 1913,8 by the United States Bureau of Mines in 1938,4 by Hancock in 1947,3 and by the national French research agency Cerchar in recent years.2 In 1950 Anderson' proposed the use of this criterion as an efficiency value to replace the Frazer and Yancey formula. However, none of the above-mentioned investigators used the misplaced material concept in a manner that would provide universal coal-cleaning equipment comparisons. The Correlation Theory The ideal coal cleaning process would treat all sizes and would make a perfect separation at any given specific gravity. All material lower in density than the desired value would report in the coal product and all material higher in density would report in the refuse product. Unfortunately, no known cleaning process achieves this goal and there seems little likelihood that any process yet to be invented will do more than approach it. When coal is treated in volume under operating conditions, it is impossible to avoid mechanical entrapment, fluctuations in throughput and effective gravity of separation, and the creation of turbulent currents, even when a true heavy-liquid bath is used and the feed is closely sized and contains little intermediate gravity material. This being so, it is possible to appreciate the difficulties inherent in trying to obtain a perfect separation when treating a wide range of sizes and a feed containing high percentages of intermediate material, using turbulent currents to help create the effective separation gravity, under operating conditions which normally tend to be on the overload side. When coal is separated from refuse in any coal cleaning equipment, some refuse always reports to the coal and some coal to the refuse; the writer therefore assumed that there should be a relationship between the total amount of misplaced material produced by any given piece of equipment and the difficulty of separation as represented by the percentage of near gravity material in the feed. With small amounts of near gravity or ±0.1 material in the feed there should be less misplacement of material than would occur with large amounts of near

Jan 1, 1952

By E. P. Sadowski, L. P. Trudeau, C. R. Cupp

The apparent high radiation resistance of two varieties of maraging steel is described and an indication is given of some phenomena that require further study. Two aspects were included in this work: a) effects of neutron irradiation on the room-ternperature tensile properties of 18 Ni (250) maraging steel; b) effects of neutron irradiation on the room-temperature tensile properties and on the impact resistance of 12 pct Ni-5pct Cr-3 pct Mo maraging steel. Subsize tensile and impact test specimens were used. The 18 pct Ni steel was irradiated to a neutron fluence of 3.3 x 10 n per sq cm (>1 mev). The 12-5-3 variety was irradiated to two neutron fluence levels: 1) 1.5 x 10 per sq CTTL (>I mev) 2) 3.1 x 10'9nper sq cm (>1 mev) In all respects both varieties showed unusually high resistance to change in strength and ductility after neutron irradiation. There was no noteuorthy over-aging in fully aged specimens. The aged 18 pct Ni variety showed a 4 pct drop in yield strength upon irradiation with little or no decrease in ductility. The aged 12-5-3 variety showed a 5 pct increase in yield strength on irradiation with a very slight decrease in ductility. The impact transition of the 12-5-3 variety only increased 35 to 45 Co upon irradiation, and the shear energy fracture value for the subsized specimens decreased only half as much as expected frorn experience with conventional steels. THE development of new reactor concepts and nuclear technology now depends to a large extent on the availability of new alloys capable of operating in stringent reactor environments. One of the most important new types of metallurgical material that has been developed in recent years is the maraging steels. This type of alloy is notable for its remarkable retention of toughness at high strength levels. In the course of a research program which incorporated ambient temperature radiation effects studies to determine the mechanisms of strengthening of 18 pct nickel maraging steel,' it became evident that the resistance of the alloy to neutron-induced changes in mechanical properties should be studied. This paper is to present evidence for the apparent high radiation resistance of maraging steel. The experimental program was in two parts: a) Effect of neutron irradiation on the room-temperature tensile properties of 18 Ni (250) maraging steel. b) Effect of neutron irradiation on the room-temperature tensile properties and on the impact resistance of 12 pct Ni-5 pct Cr-3 pct Mo maraging steel, a cobalt-free alloy. EXPERIMENTAL PROCEDURE 18 Ni (250) Maraging Steel. Tensile test specimens were machined to the dimensions shown in Fig. 1 from $-in.-diam hot-rolled bar of the composition shown in Table I. All specimens were austenitized for 1 hr at 1500°F (815°C) in argon and water-quenched. Half of the specimens were then mar aged for 3 hr at 900° F (482°C) in argon. The specimens were irradiated in a fast neutron transformer rod at approximately 120°F (50°C) to a neutron fluence of 3.3 x 10 10 n per sq cm (>1 mev). The irradiation was performed in the NRX reactor of Atomic Energy of Canada, Ltd., at Chalk River, Ontario, Canada. In all cases, the thermal neutron flux was three times as high as the reported (> 1 mev) fluxes. The room-temperature tensile tests were performed with an Instron tensile testing machine operating at a crosshead speed of 0.05 in. per min. The loading frame of the machine, mounted in a hot cell, was remotely controlled. The specimens were tested either

Jan 1, 1968

By M. R. J. Wyllie, Walter D. Rose

The use of electrical well logs for the quantitative determination of such reservoir parameters as connate water saturation, formation permeability and connate water salinity has recently been attracting the attention of a number of workers1,2,3,4,5. While the theory of the determination of connate water salinity from the self potential S.P. log has received rather detailed treatment, relatively little attention has been paid to the theoretical aspects involved in the quantitative interpretation of resistivity data. It is clear that if electrical well logs can be used for the quantitative evaluation of physical characteristics of reservoir rock, they will provide a valuable tool to supplement cheaply information obtained by more laborious core analysis. In certain cases it is conceivable that the coring program could be considerably curtailed if the electric log could be relied upon to give reasonably accurate quantitative information. It is our object in this paper to examine the theoretical basis of quantitative log interpretation as expressed in such well-established logging concepts as formation factor and cementation factor. It is also our object to investigate the physical aspects of the rela- tionship which is presumed to exist between resistivity index and brine saturation in reservoir rock. In particular, we will endeavor to draw attention to the fact that it is possible to express these logging concepts in terms of capillary pressure-saturation relationships, permeability and tortuosity, parameters which we will consider in this paper to be fundamentally indicative of rock texture. The probability of being able to obtain from log recordings alone the data theoretically essential to permit quantitative log interpretatiou will be examined, and consideration will also be given to the problem of formulating simple semi-empirical relationships for use in the field. THEORETICAL CONSIDERATIONS The concept of formation resistivity factor, or as it is now commonly called, formation factor, appears to have been introduced by G. E. Archie",". Formation resistivity factor as defined by Archie is the resistivity of a rock 100 per cent brine-saturated divided by the resistivity of the brine. This relationship had previously been used by physical chemists and the concept is, for example, implicit in an early treatment by Fricke7 of the conductivity of aqueous slurries. Archie was able to establish the fact that although the relationship between formation factor and porosity did not change too greatly for different rock textures, the relationship between formation factor and permeability was markedly dependent on rock texture. The relationship found by Archie between formation factor and porosity was expanded by H. Guyod8. Archie had suggested that the formation factor, F, could be expressed in terms of the porosity, F, as follows: F =F-m .... (11 where, F = effective porosity fraction of the sand, and, m = an exponent greater than about 1.3. Guyod introduced the term cementation factor for the exponent, m, since it may be qualitatively shown that formation factor for any given porosity tends to increase as the sand becomes more cemented. It would appear that the concept of m as an exponent representing the cementation in a sand is strictly only of qualitative value, since the quantitative estimation of m by means other than the determination of F and F has not been possible. Guyod suggests that m can be estimated from a visual examination of a sample of rock, but this method appears to be both arbitrary and unreliable. It can be said, however, that in intergranular reservoirs of practical interest, the range of m is from about 1.4 to 3.0. A possible explanation for the fact that m is found to lie between these limits will be offered below. The determination of formation factor is an essential preliminary to the

Jan 1, 1950

By T. M. Geffen, J. L. Sanderlin, F. F. Craig, D. W. Moore

Scaled reservoir models have been used to study the effect of gravity on oil recovery performance in frontal-drive operations; namely, water, gas, or solvent flooding. The difference in density between the reservoir oil and the injected fluid causes their segregation, resulting in a non-uniform advance of the fluid front. In the laboratory flow tests, which simulated both five-spot and linear injection operations in flat reservoirs, the viscous, capillary, and gravity forces present in these operations were scaled. Dyed fluids were used so that the gross movement of the injected fluid could be observed. The studies covered a range of injection rates, formation thicknesses, and rock and fluid properties normally encountered in field operations. The results of the model tests indicate that the volume of the reservoir contacted by the injected fluid at its breakthrough into the producing well is less than that expected based on information which neglects gravity effects. This difference can often be as much as 80 per cent by gas or water injection in uniform sand bodies. Preliminary flow tests on a non-uniform sand body indicate that the uniformity of the flood fronts may in some situations be influenced to a much greater degree by permeability variations within the rock body than by gravity effects. The magnitude of fluid segregation due to gravity is controlled by the average injection rate, rather than day to day or week to week variations. INTRODUCTION One of the important factors controlling the oil recovery from a frontal-drive operation, such as water, gas, or solvent injection, is the volumetric sweep efficiency. This factor is a measure of the gross portion of the reservoir that is contacted by the displacing fluid. The volumetric sweep efficiency is influenced by gravity effects, well arrangements, and variations in rock permeability within the reservoir. The gravity effects are due to the displacing fluid being of different density than the reservoir oil. This causes the displacing fluid to move preferentially toward either the top or bottom of the formation. Non-uniform advance of the flood front can be caused by the relative positions of the injection and production wells. Variations in rock permeability also result in an uneven advance of the displacing fluid. Gas zones, representing paths of low resistance, may result in non-uniform movement of the injected fluid. The more uneven the advance of the displacing fluid, the lower is the volumetric sweep efficiency and thus the lower the oil recovery efficiency. The effects of well arrangement and permeability variations, upon the gross movement of the injected fluid have been the subjects of extensive investiga-ti0ns.1,2,3,4,5,6 Gravity effects have been recognized as being present in frontal-drive operations'.' but there has been little quantitative measure of their magnitude. Very likely the reason gravity effects have not been studied is that this problem does not lend itself to simple experimental or mathematical analysis. The development of reservoir modeling techniques has made possible a quantitative determination of the effect of gravity on fluid segregation and thus upon the oil recovery performance of frontal-drive operations. This paper presents a progress report on a laboratory investigation of gravity effects in fully liquid-saturated, horizontal, uniform as well as non-uniform systems. These results were obtained from scaled experiments on both linear and five-spot systems within the range of conditions normally encountered in secondary recovery operations. This study is concerned with the gross movement of the injected fluid to the time it first reaches the production wellbore. Therefore, it should be emphasized that the results of this paper can not be used to determine the effect of gravity segregation on oil recovery efficiency at abandonment conditions. MODEL SCALING The science of scaling models is relatively new as applied as petroleum production problems, Many of the scaling techniques have been adaptions of those used in basic studies of heat and fluid flow. There are

Jan 1, 1958

By R. H. McLean

Velocity, kinetic energy and shear in crossflow beneath three-cone jet bits may influence cleaning of the bottom of the borehole and the teeth of the bit. Laboratory investigation shows that each of these parameters is a function of the diameter of the borehole and the product of the volume rate of flow and velocity through the nozzles (QVn). Increasing QV, or decreasing the diameter of the borehole increases each parameter. These functions provide means of predicting the magnitude of each parameter and of scaling the cleaning forces. INTRODUCTION In drilling operations using conventional jet-type rock bits, the impinging jels create an important flow mechanism. Called crossflow, this flow mechanism originates in the impact area of the jets, spreads across the bottom of the hole and supplies the principal source of energy to clean the teeth of the bit and most of the bottom.' Besides providing the means of cleaning cuttings from the bit and the hole bottom, crossflow may also have other, less direct, effects on the rate of penetration. The shear stress generated on the bottom by crossflow will inIluence the thickness and permeability of any filter cake of mud solids or crushed material which forms on the bottom.' These factors may affect the rate of penetration. A previous pulblication introduced some fundamental concepts of crossflow.' Crossllow was shown to occur in a thin layer adjacent to the bottom, and to cover the bottom completely. The maximum velocity in the crossflow above any position on the bottom was found to be directly proporlional to the square root of the product of the volume rate of flow and velocity through the nozzles—the jct QV,,—and inversely proportional to the diameter of the borehole. This information indicated that the effectiveness of crossllow in scavenging the bottom can be improved by maximizing the jet QV. The investigation reported herein amplifies the definition of crossflow. Complete velocity profile data above a representative position on the bottom are analyzed. These data better illustrate control of the capacity of crossflow to scavenge the bottom, and also relate shear stress on the bottom to known, controllable parameters. The conclusion reached in the previous publication that maximization of the jet QV. produces the maximunl cleaning beneath current jet bits is unchanged by these new data; rather, it is strcngthened. Data presented here show that the kinetic energy flux above a representative position on the bottom is maximized by maximizing the jet QV. The shear stress on the bottom will also be shown to be maximized in the same manner. Since the functions relating the jet QV. to velocity, shear stress and kinetic energy also involve the diameter of the borehole, means of equating, or scaling, these quantities in different sizes of boreholes will be illustrated. EXPERIMENTAL EQUIPMENT AND TECHNIQUE JET BIT MODEL Data were recorded from the same laboratory model as used in the aforementioned investigation of the flow around a jet bit.' The model consisted of a 43/4-in. three-cone, jet-type rock bit in a smooth, flat-bottomed borehole constructed of lucite. The bit had a shape and nozzle placement closely resembling larger three-cone jet bits commonly used in field operations. Fig. 1 illustrates the impact area on the bottom of a jet from this bit. Details of the orientation of the jets may be found in the previous publication. TECHNIOU'E OF MEASUREMENT Measurements of the crossflow were made by inserting a very small Pitot tube through the bottom of the simulatetl boreholc. Extreme thinness of the layer of crossflow necessitated accurate measurements of the height of the Pitot tube above the bottom to achieve close definition of the velocity profile. A cathetometer, which could be read to the nearest 0.005 cm, was used to make this measurement.

Jan 1, 1966

By C. S. Matthews, P. Hazebroek, F. Brons

A method has been developed for calculating the average pressure in a bounded reservoir. The reservoir is first divided into the individual drainage volumes of each well, by using the criterion that at steady state each individual drainage volume is proportional to a well's production rate. The average pressure in each drainage volume is then calculated by a method developed in the report. By volumetrically averaging these individual drainage volume pressures, the average pressure in the entire reservoir is obtained. To calculate the average pressure in each drainage volume, a correction is applied to the ordinary extrapolated pressure, i.e., the pressure obtained by extrapolating to infinite time the linear portion of the graph of closed-in pressure versus log[?t/(t + ?t)], where At is the closed-in time and t the production time. The correction, which is a function of the production time, is presented in graphical form for different shapes of the drainage area (horizontal cross section of the drainage volume). INTRODUCTION It is important to be able to find the volumetric average pressure in a reservoir so that the size of the reservoir may be determined from material balance calculations. It is also desirable to be able to find the approximate distribution of pressure within a reservoir for detection of fluid movement. 'The purpose of this paper is to present a method for calculating both the average reservoir pressure and the approximate distribution of pressure within a bounded reservoir — that is, a reservoir with no water drive. In reservoirs where the pressure builds up rapidly after wells are shut in, the determination of average pressure generally poses little problem, for one often need only average the final buildup pressures. It is when pressure buildup is slow that difficulties arise. For practical and economical reasons, the time allowable for closing in wells is limited. If at the maximum allowable closed-in time the pressure has not reached a con- stant value (and this is more often the case than is generally realized), calculation of average pressure presents difficulties. One method which has been used in the past for obtaining an "average" pressure involves the extrapolation to infinite closed-in time of curves of pressure as a function of closed-in time. Contour maps of these extrapolated pressures are used in arriving at an average pressure. In some cases wells are shut in for arbitrary Lengths of time, and contours of these arbitrary shut-in pressures are used for calculation. All these methods are somewhat arbitrary and suffer from lack of fundamental theoretical basis. For example, no precise meaning can be given either to the extrapolated or to the arbitrary shut-in pressure. Furthermore, there seems to be no reason why either set of pressures should be contourable. It was because of the vagueness of these old methods and their lack of a sound theoretical basis that, the present method for determining average reservoir pressure was developed. SYMBOLS A = drainage area of a well, cm' A, = drainage area of well j, cm2 A , = total drainage area of all wells in a reservoir, cm2 c = coefficient of compressibility of fluid at reservoir temperature, atm-1 f = hydrocarbin-filled porosity, fraction h = net thickness of formation, cm k = permeability of formation, darcy p = pressure at a well, atm p0 = initial reservoir pressure, atm p* = pressure obtained by straight-line extrapolation of linear portion of plot of p versus log [At/(t+ At)] p = volumetric average pressure inside drainage volume of a well q = volume flow rate at a well, cc/sec a: prevailing reservoir conditions q, = total volumetric flow rate from a reservoir, cc/sec at prevailing reservoir conditions rw = radius of well bore, cm T = kt/fµcA t = corrected time of production of a well, seconds, defined by

Jan 1, 1955

By Nigel B. Pearson

INTRODUCTION Chingola Division of Nchanga Consolidated Copper Mines Limited is situated in the Copper- belt of Zambia, approximately 30 kilometres south of the border with Zaire. It is the country's leading producer of copper and its principal foreign exchange earner, and has been, in its time, the largest single copper producer in the British Commonwealth and the second largest in the world. There are two 'primary' copper sources at the Division, namely Underground producing approximately 100 000 tonnes of contained copper annually, and the Open Pits, producing approximately 200 000 tonnes annually. Following the commissioning of the Tailings Leach Plant in 1974, treating final tailings, metallurgical recovery increased from approximately 70% to in excess of 80%. Extensive deposits of cobalt are present and following the rise in the market price in 1979 and 1980, plans were made to extract this separately. Cobalt production commenced in 1980 from Underground and will increase substantially when the Open Pit cobalt reserves become available in approximately 2 years time. This paper covers the development of the production method of the principal Underground Orebody, known as the Lower Orebody, and the control systems applied to it. HISTORY The Lower Orebody, formerly known as the Nchanga West Orebody, was discovered by extensive Diamond Drilling operations in the late 1920's. Development of the Shaft complex and the orebody commenced and continued until 1931 when flooding from development on the 480 foot horizon necessitated closure. The mine remained closed until 1937 and production commenced from the Lower Orebody in 1939. Various expansion schemes of the Metallurgical Plants and the Underground workings were completed over the years with the result that Underground production increased steadily to in excess of 150 000 tonnes copper annually in the 1960's. In recent years production has levelled out at slightly less than 100 000 tonnes annually. This decrease is as a result of producing from a lower grade reserve, available within a more limited production front, and the application of production constraints in the light of experience gained. GEOLOGY Within the Nchanga Mining Licence Area of Chingola Division, there are a total of 14 known copper orebodies in the basal part of a thick succession of sediments of Pre-Cambrian age. The sediments rest on a Basement Complex of granites, gneisses and schists. Three superimposed stratiform orebodies, known locally as the Lower, Intermediate and Upper Orebodies, occur over a vertical interval of approximately 150m. Current annual copper production is approximately 300 000 tonnes. The ore originates from four different mining operations, one Underground mine and three Open Pits. Approximately 9 million tonnes of ore grading between 3 and 3.5% copper are milled annually. The two largest sources of ore, the Under- ground mine and the Nchanga Open Pit are on the gently dipping south limb of the Nchanga Syncline (Figure 1). In the Nchanga Open Pit the major source of ore is the Upper Orebody with lesser tonnages from the Intermediate and Lower Orebodies. In the Underground mine the Lower Orebody is the only orebody being mined, underground mining of the Upper Orebody being suspended for economic reasons in 1978. The typical stratigraphic column is illust- rated in the legend on Figure 1. The footwall rocks comprise either Nchanga Red Granite or the Arkose formation, both of which are hard competent rocks, offering generally good mining conditions for footwall development. Copper mineralisation of the Lower Orebody extends from the upper part of the Arkose to the top of the Lower Banded Shales. Overlying the Arkose is the Transition zone (0-18111 in thickness). The Transition zone comprise a lower member, Transitional Arkose, consisting of incoherent kaolinised sandstone, and an upper member, Transitional Lower Banded Shales consisting of normally soft bleached and silicified shale. The soft nature of the Transitional rocks is

Jan 1, 1981

By W. A. Tiller, S. O’Hara

Grain refinement in stirred melts has previously been shown to arise from dendrite segmentation. The present work discusses experiments capable of distinguishing between remelting and mechanical effects which result from fluid motion. Nucleation of still un-dercooled Sn-1 pct Pb melts with a single-crystal seed shows that the grain size of the resulting ingots decreases as the undercooling at nucleation increases up to about 1Z0C, then increases again for larger undercoolings. Pure melts exhibit little refinement as a result of nucleation at any undercooling up to 12°C. The presence of fluid motion gives rise to grain multiplication in pure melts, but the magnitude of this effect is partly obscured by solid-state pain growth. The results indicate that two grain-multiplication pvocesses can operate, one which involves dendrite remelt ing, and another associated with fluid motion. Small ingot experiments indicated that an equiaxed transition can be caused by fluid motion. THE presence of forced convection during recales-cence of undercooled Sn-1.3 pct Pb melts has been shown to give rise to extensive grain refinement during solidification of small ingots of this alloy.''2 In Ref. 2, this refining was shown to result entirely from dendrite fragmentation and not from any nucleation mechanism. The purpose of the present work is to establish what processes lead to dendrite segmentation. Experiments were therefore carried out to determine the resulting structure of both pure tin and Sn-1 pct Pb melts, nucleated at various undercoolings by single-crystal seeds in the absence of forced or natural convection. By comparing the information obtained from these experiments with the results of Refs. 1 and 2, the effects of fluid motion on the ingot structure, and the details of the segmentation mechanisms, should become clear. The results of these experiments should also be helpful in understanding the solidification structures of larger ingots since natural convection, which is known to play an important role in determining ingot structure13 differs only in degree from the fluid-motion effects studied here. I) EXPERIMENTAL PROCEDURE The apparatus shown in Fig. 1 was constructed to permit controlled nucleation of stirred or unstirred supercooled melts using a seed oriented in the tin dendrite growth direction. Because the melts, in all cases, were capable of larger supercoolings than the one used in each experiment, the structure observed always resulted from a single nucleation event at the surface. The sequential procedure was as follows: Tin was melted under purified hydrogen and the oxide removed at a temperature of 450°C . Electromagnetic stirring was then used to clear the surface completely and during this time the guide tube was lowered underneath the melt surface. A pointed Czochralski-like crystal was then propagated from a specially prepared tin dendrite seed4 inside the guide tube while the temperature of the bulk melt was held just above the melting point, ensuring a perfectly clean seed which was epitaxial with the original dendrite. The dendrite and epitaxial seed, still in the guide tube, were withdrawn from the melt which, in turn, was undercooled to 2C° below the temperature at which the nucleation experiment was to be performed. The melt was held there for several minutes, both in stirred and in still-melt experiments. The temperature was next raised to the required value and the melt nucleated by the carrot seed inside the guide tube, the heater power being shut off just after nucleation. In experiments where quenching after nucleation was required, a high-density graphite crucible was used, but otherwise the procedure was as described above. The temperature gradient in the melt before nucleation was kept below 1C° per cm by carrying the experiment out just below the center of the heater shown in Fig. 1.

Jan 1, 1968

By C. R. Masson, M. L. Pearce

The nitrogen contents of seven specimens of iron and steel were determined by the three methods. Good agreement won generally observed between the results of the isotope -dilution and Kjeldahl determinations, tltzts establishing the reliability of the isotope -dilutzon technique. For some specimens, the vacuum-fusion values at 1650 °C were lower than the corresponding Kjeldahl values by up to 0.0015 pct absolute. These differences were eliminated for a specimen of "pure" iron by performing the vacuum-fusion analyses at 2100 ° to 2240 °C. The reason for this relatively high temperature required m the vaczium-fusion analyses is discussed. ThE determination of nitrogen in iron and steel has been the subject of numerous investigations and a concise review of the extensive literature in this field has recently appeared.' Two methods of analysis are most commonly employed, the vacuum -fusion and the chemical or Kjeldahl method. Except in special cases where the presence of certain alloying elements may present difficulties, both methods are generally regarded as applicable, although the vacuum-fusion method is more sensitive and is often preferred for samples of very low-nitrogen content. The method of isotope dilution, first employed for the determination of gases in metals by Kirshenbaum and Grosse,2 has recently been applied to the determination of nitrogen in iron and steel.3,4 The method is attractive in that it does not require the quantitative extraction of gas from the metal specimen. Staley and Svec3 modified the standard Kjeldahlpro-cedure by introducing, along with the steel sample, a measured amount of ammonium sulfate enriched in the isotope 15N. The nitrogen content was obtained by measuring the isotopic ratio 15N/14N in the resultant solution. The technique involved the distillation of nitrogen as ammonia from alkaline solution and subsequent oxidation of NH4+ in the distillate to gaseous nitrogen. The results for iron and steel samples of commercial origin were often in reasonable agreement with values quoted by the suppliers, although occasional marked discrepancies were observed, particularly with the results of vacuum-fusion analyses. Preliminary work in this laboratory4 on the determination of nitrogen by isotope-dilution in the gas phase yielded results for three samples which were significantly higher than the values obtained by vacuum-fusion. The present paper describes an extension of this work to a wider variety of samples. Analytical results by the Kjeldahl method are also reported. EXPERIMENTAL The technique employed for the isotope-dilution determinations was similar to that described elsewhere.4 Several minor modifications were made to the apparatus. These included the use of molybdenum instead of platinum-rhodium suspending wires when crucibles were degassed at temperatures above 1600°C. The circulatory system used previously for converting CO to CO2 was replaced by a single chamber, the walls of which were cooled directly in liquid nitrogen. Temperatures were measured with an optical pyrometer which had been recently calibrated and, as in the previous study, were minimum values obtained by sighting on the center bottom of the crucible. Temperatures obtained by sighting on the interior walls were always higher by 100' to 130°C with the design of crucible employed. Flat-bottomed crucibles exhibited an even larger temperature difference. To check the efficiency with which crucibles were degassed in the isotope-dilution apparatus, several experiments were done in the vacuum-fusion apparatus, suitably modified4 for isotope-dilution studies. Faster pumping speeds and higher degassing temperatures were possible, the latter due to the use of a generator with a higher power output (10 Kva at 398 kc per sec). In these experiments the combustion of CO was omitted and a correction4 was made for the contribution of CO to the ion current at m/e 28, used in determining the isotopic ratio R = l5NP4N. This correction did not seriously influence the re-

Jan 1, 1962

By Charles E. Mann, Douglas J. Lootens, Rufino B. Bomasang

INTRODUCTION The economy of the Philippines, like that of many other developing nations, has been severely strained by the rapid escalation of imported crude oil prices over the past decade. In an effort to reduce its dependence on imported crude, significantly increased emphasis, both in the public and private sectors, is being placed on the rapid development of Philippine coal resources. This paper is designed to present an overview of the Philippine coal industry, its historical development, present status, and future potential. COAL RESOURCES AND RESERVES In 1977, a British consulting firm collated all available data on Philippine coal resources and estimated the total coals potential at about 0.9 to 1.4 billion metric tons (mt). In 1981, using more recent data the Philippine Bureau of Energy Development (BED) estimated the potential at 1.7 billion mt. The BED estimated the proven coal reserves at 257 million mt. These coal resources are scattered throughout the major islands comprising the Philippine archipelago. However, about 52 percent of the currently proven reserves are located in the tiny island of Semirara between the islands of Mindoro and Panay. About 25 percent is located in the Cagayan Valley in northeastern Luzon while 13 percent is located in Mindanao. The remaining 10 percent is scattered in the islands of Cebu, Samar, Mindoro, Negros, Polillo, Batan, and Catanduanes. The bulk of Philippine coal is sub-bituminous in rank. Some bituminous coal with coking properties has been reported in Malangas, Zamboanga del Sur in western Mindanao, and in Catanduanes, but the exact tonnages have not been accurately established. It is believed, however, to be at most 5 million tons. A little semi-anthracite coal exists in Malangas with the remaining unmined reserves estimated at a little more than one million tons. Table 1 shows the quality ranges of Philippine coal deposits.. Figure 1 shows the distribution of known Philippine coal resources PHILIPPINE COAL PRODUCTION In 1981, total coal production was 330,720 mt. Though modest by world standards, this was the highest production figure in Philippine coal mining history. This came from 25 mines, mainly small scale underground mines of capacities varying from 10 to 100 tons per day (tpd). The actual production capability was about 800,000 mt in 1981, but the producers had to scale down production to the level of local demand. Even then, some 50,000 tons of coal remained unused by year-end. The 1981 production came from the following areas: Cebu 224,751 mt (68%) Zamboanga del Sur 57,095 mt (17%) Semirara 13,222 mt (4%) Polillo 17,132 mt (5%) Cebu has always been the main source of Philippine coal production. Coal mining has been carried out in Cebu for the last 50 years, although it has been characterized by a very crude mining technology. Due to the limited size of coal deposits and complicated geology, however, Cebu's share of the total national production is expected to go down to only 10 percent by 1985. Except for Semirara's open-pit production, the balance of Philippine coal production comes from underground mines. These underground mines generally use room-and-pillar methods, except for those at Zamboanga del Sur and Polillo which use modified longwall methods patterned after those used in Taiwan. Due to the much higher mining recovery, most underground mines are now planning on switching to longwall. At Semirara, an open pit mine is now being developed for a one million ton per year capacity by 1984, and will then be the largest Philippine coal mine. While present earthmoving work is done with bulldozers, loaders, and trucks, the main coal extraction equipment will be bucket wheel excavators imported from Austria. PRESENT COAL UTILIZATION Existing coal production is sold mainly to coal- fired cement plants and power plants. The balance is sold to a sugar central, a coconut oil mill, a food processing plant, and two plants producing carbide and ferro-silicon alloys.

Jan 1, 1982

By H. Majima, E. Peters

The oxidation rates of pyrite, pyrrhotite, chalcopy-rite, chalcocite, covellite, bortzite, galem, sphalerite, and stibnite have beet2 carefully compared at 120 oC, using aqueous phosphate solutions buffered at pH 2.7, 7.1, and 11.2 as zvell as in 1.0 M NaOH and in am~onia-containing solutions. Under conditions of these studies, variations in rates of more than a factor of two per unit of mineral surface area were not observed for different minerals in phosphate buffered solutions, except that pyrrhotite showed unusually rapid oxidation rates In acid solutions, and stibnite showed very rapid oxidation rates in basic solutions. In caustic solutions, oxidation vates varied considerably, the most rapid being for pyrite and the slowest for sphalerite. In the presence of ammonia, only copper minerals showed more vapid oxidation rates in neutral solutions. Oxidation rates were measured on the basis of oxygen-consunzption rates per unit surface area of exposed mineral, using equivalent surface areas, mineral mesh sizes, pulp densities, agitation, and oxygen pressure. OXIDIZING pressure leaching of sulfide minerals has been a commercial method of treating concentrates of nickel, cobalt, and copper for over a decade, and recently the treatment of lead and zinc concentrates,' and even the recovery of sulfur from pyrrhotite by such processes, have been shown to be technically feasible.' Furthermore, the leaching of uranium ores by solutions acidified through the oxidation of pyrite and other sulfides present in the ore is commercially practicable.3 5 The mechanisms by which sulfide minerals oxidize in aqueous environments are not yet understood. However, certain generalities can be made for systems containing no unusual reagents, in which oxygen is the oxidizing agent. In acid solutions the oxidation of many sulfide minerals leads to the formation of elemental sulfur.8'7 The attack of some sulfides, notably galena (PbS), sphalerite (ZnS), and pyrrhotite (FeS), at temperatures below about 110°C leads to an almost quantitative conversion of the sulfide mineral to elemental sulfur according to the equation: On the other hand pyrite (FeS2) and copper minerals do not yield elemental sulfur quantitatively. Not more than 50 pct of the sulfur in pyrite is converted to the elemental form in acid oxidation, and this percentage is lowered by higher temperatures and possibly higher oxygen pressures.8'9 Copper minerals sometimes yield elemental sulfurlo'" but, since sulfur can reduce cupric ions with the precipitation of CuS or Cu2S, it is not likely that elemental sulfur is obtained in good yield. In acid solutions sulfur that is not converted to the elemental form always appears as sulfate salts or sulfuric acid. Sulfide minerals that yield elemental sulfur nearly quantitatively on oxidation in acid solutions tend to generate H2S when the conditions are not oxidizing. However, although it has been suggested that H2S is an essential intermediate in the mechanism of oxidation," this is doubtful because it is oxidized very slowly by oxygen. In basic solutions elemental sulfur is never formed, as predicted by thermodynamical considerations, summarized in the form of a potential-pH diagram in Fig. 1. However in neutral and basic solutions, unstable forms of sulfur not shown in the diagram appear during oxidation, notably thiosulfate (SOT-) and polythi-onates (S,O;-, x = 2 to 6).l3,l4 On prolonged oxidation only hexavalent sulfur as sulfate (SOi-) or sulfamate (NHzSO;) is found to remain, the latter appearing only in the presence of ammonia. The oxidation kinetics for thiosulfate by oxygen are the subject of a separate paper,'5 and were studied on the presumption that thiosulfate may, in fact, be a primary oxidation product for sulfide minerals, at least in neutral and basic solutions. A survey of the literature indicated that comparative oxidation rates for different sulfide minerals were needed to establish the role of the mineral structure and its thermodynamic properties on the oxidation rates and mechanisms.

Jan 1, 1967

By F. A. Schaufelberger

METAL can be recovered from a leach solution either indirectly by precipitation as a compound that is later reduced or directly by electrolysis, cementation, or chemical reduction, for example, with hydrogen. The widely used electrowinning processes suffer from the inefficiency of converting fuel to low voltage dc and from low current efficiency when complete recovery is attempted. With the batch hydrogen reduction methods outlined here, a unit of fuel converted to hydrogen in modern gas reform plants will make two to three times as much metal as can be obtained with electrolysis. With continuous hydrogen reduction, four to six times as much metal is obtained. Precipitation by reducing a metal salt solution with H, has been known for almost 100 years.' Commercial use of the process, however, awaited the research and development program initiated by Chemical Construction Corp. Within the last few years this program has brought about construction of commercial plants, listed on page 545. Ideal hydrogen reduction will precipitate a pure metal from a solution obtained by commercial leaching methods at a rapid rate without excessive temperatures and pressures. The metal precipitate will be of desired size and density, less than a percent left in the vessel being deposited on the wetted parts. The present discussion will outline Chemical Construction Corp.'s early development program and will discuss the chemistry and mechanics of reducing copper, nickel, cobalt, and cadmium from solution by H2. Work on selective reduction of nickel from cobalt has been described earlier.' The article is not concerned with precipitation of copper by disproportionation of cuprous solutions where yield is limited3 to 50 pet.* It does not discuss use of gaseous reducing agents other than hydrogen, such as SO2,4 which may lead to contamination with sulfur, or C0,5 which is considerably more expensive than hydrogen and produces a gaseous reaction product, CO2.† The article does not include use of other nongaseous reducing agents,7 which are far more expensive than CO. Also, notes on the history, engineering design, and performance of commercial plants have been given elsewhere.' It will be shown that a pure metal can be precipitated by reduction with hydrogen from solutions obtained by commercial leaching methods when the solution composition is controlled and proper acidity, proper metal ion concentration through controlled complex formation and hydrolysis, and adequate agitation for suspension of metal and for transfer of reducing gas are maintained. Reasonable temperatures and pressures are used which are less excessive than those used by previous workers, although they are maintained at above equilibrium values because of process kinetics.' It has also been found necessary to induce nuclea-tion and to control growth and agglomeration in order to get a powder product of suitable physical properties. Review of the Literature Muller, Schlecht, and Schubardt of I. G. Farben10 claimed a successive reduction of silver, copper, nickel, cobalt, and zinc from ammoniacal solution by applying progressively higher temperatures and hydrogen pressures. Metals produced by this technique were not pure, since no attempt was made to adjust the solution composition between the different reductions. It is doubtful that the zinc product contained any metallic zinc. The major contribution to the literature" was made by the Ipatiews,12 who prepared platinum, iridium, copper, nickel, cobalt, lead, tin, arsenic, antimony, and bismuth at often extreme conditions, up to the critical temperature of water (373ºC), at 1500 to 7500 psi, for as long as several days. Even cadmium and zinc were claimed to have been observed in trace quantities precipitated together with basic salts. In the early experiments solutions were not stirred. Extensive formation of oxides and basic salts occurred, probably long before sufficient hydrogen could be supplied in the unstirred liquid to decrease the metal concentration in solution by reduction. The importance of hydrogen pressure was

Jan 1, 1957

By P. L. Raffo, T. E. Mitchell

Single crystals of Ta-Re alloys have been deformed in tension and compression as a function of solute concentration and temperature. The temperature dependence of the yield stress is decreased by alloying. At low temperatures there is a minimum in the yield stress (at about 2 at. pct) as a function of concentration. Three-stage hardening was not observed in the alloys. Linear hardening rates, where observed, were characteristic of stage II in pure tantalum. At low temperatures and high concentrations, work softening was often prominent at high strains. Electron microscopy revealed that dislocations are mostly of screw orientation, although heavily jogged and cusped, and often occurring in dipole pairs. The alloys become more prone to cleavage with increasing rhenium content, due mostly to an increasing yield stress rather than a change in cleavage strength. Alloying apparently decreases the Peierls stress of bcc metals at low temperatures by facilitating kink nucleation at solule atoms on screw dislocations. Higher solute concentrations further increase the yield stress. Quantitative calculations and comparisons with other bcc alloy systems are made to show that the predominant hardening mechanism is due to the interaction of dislocations with misfitting solute atoms close to their slip planes. The interaction is so localized that the core dilatation of dislocations could be important. CONSIDERABLE interest has arisen in recent years concerning the mechanical properties of unalloyed bcc metal single crystals.'-7 These and similar studies on polycrystalline materials have resulted in a considerable understanding of the macroscopic yielding characteristics8, 9 and the process of work hardening.'' This latter advance is due in a large part to the increased use of transmission electron microscopy in P I RAFFO junior Membar AlMF is Metallyureist Iewis such studies and the general observance of three stages of work hardening in the shear stress-shear strain curves. The properties of substitutional alloy single crystals of bcc metals have not been determined in as much detail as in the pure metals. Insufficient systematic studies have been made on the effect of alloying on the magnitude and temperature dependence of the yield stress or the stress-strain curves. The alloy systems which have been studied to some extent have been Ta-W,11, 12 Ta-Mo," Ta-Nb,12'14 Mo-Re,15 W-Re,16 and alloys of iron with nickel,17 platinum,17 vanadium,18 and aluminum.18 A large body of data exists on Fe-Si alloys, but these have mainly been a result of the experimental convenience of using this material in place of unalloyed iron, rather than a comprehensive study of the mechanical properties of the alloy system per se. The present authors have previously reported additional data on the properties of Ta-Re, Ta-W, and Ta-Mo alloy single crystals.1g Some of the data on Ta-Re alloys will be described later in this paper. Briefly, single crystals of Ta-Re (1.2 to 10.6 pct), Ta-W (6.2 and 10.2 pct), and Ta-Mo (2.9 and 9.0 pct) were grown by the floating zone method in vacuum. Compression tests were performed in the temperature range 77" to 623°K. Slip was found to occur in Ta-Re alloys at 77' and 298°K on the (~ll)[lil] system and slip lines were coarser and straighter than in unalloyed tantalum. No three-stage hardening was observed in any of the alloy systems and the curves were nearly parabolic. At temperatures above 175°K the yield stress* increased monotonically with in- creasing composition while at 77" and 175°K the yield stress initially passed through a minimum after which it began to increase again for concentrations greater than about 2 pct Re. Similar minima were observed in Ta-Mo alloys and to a lesser extent in Ta-W alloys. The minimum was interpreted as being a result of a

Jan 1, 1969

By F. L. Toy, C. E. Sims

PNEUMATIC processes for converting molten pig iron to steel were the major producers of steel during the latter half of the 19th Century and until shortly after the turn of the century, when these processes began losing ground to their rival—the basic open hearth process. The first pneumatic process for converting molten pig iron to steel was introduced almost simultaneously by Sir Henry Bessemer in England and by William Kelley in the United States during the decade 1850-1860, but it was Bessemer who pushed it to a successful commercial operation. The process itself is so basically simple that it has not lent itself to many extensive or revolutionary improvements since its inception. During the first 30 yr, numerous' designs of vessel and many modifications of practice were tried before the industry standardized on something resembling current practice. The evolution in construction has resulted in proportions and arrangements based more on mechanical expedience than on metallurgical requirements. Many of the early converters were stationary and had to be tapped similarly to the cupola or open hearth furnace. Such types were used for many years in Sweden and Germany, but the rotating or tilting types were favored in England and the United States. There was evident trouble with bottom tuyeres, becAuse about half of the early designs were constructed for side blowing with tuyeres submerged below the metal-bath surface. Even portable tuyeres, introduced from above, were tried. The principal advance, which made bottom blowing practical, was the invention, in 1863, by Henry Bessemer, and subsequent improvement by A. L. Holley, of a detachable bottom so that the tuyere section could be renewed. In the meantime, it was the demonstration, by Robert Mushet, of the value of manganese in overcoming hot-shortness in converter steel, that made the process metallurgically feasible. At first, only the acid practice was used, and Swedish pig iron, because of its low sulphur and phosphorus, was in great demand. The development of a suitable basic lining, by Sidney G. Thomas and Percy G. Gilchrist in 1876-1879, made the process applicable to the high-phosphorus ores of Europe. From the beginning of the use of the Thomas-Gilchrist process, trouble was experienced in obtaining sufficiently high temperatures in the blown metal, particularly in small heats. It is necessary, of course, to have the low-carbon blown metal at a considerably higher temperature for handling from vessel to ladle to molds than is needed for the molten pig iron. All of the increase in temperature in a pneumatic process must be obtained from the heat of combustion of the elements contained in the pig iron plus that from silicon when ferrosilicon is added. In bottom blowing practice, carbon is almost completely burned, but inasmuch as it cannot be oxidized to a higher state of oxidation than to CO when in contact with liquid iron, only a minor portion of the heat-producing capacity is utilized. Under such conditions, it is not considered to be an efficient source of heat. In the acid practice, therefore, silicon and manganese are the large heat producers, are quickly oxidized, and are efficient fuels. Too much manganese, however, thins the slag greatly and tends to promote excessive slopping or throwing of slag and metal from the converter during the blow. Most operators, therefore, like to keep the manganese content of the iron below 0.60 pct, and thus the main source of heat is the silicon, which must be at least 1 pct and sometimes 2 to 3 pct for small blows such as for castings. Ferrosilicon is sometimes added during the blow to obtain an adequate temperature level. In the basic (Thomas) practice, even more heat is required, because cold lime is added to flux the phosphorus after it is oxidized, and the basic slag must be maintained at a proper temperature. Silicon

Jan 1, 1951

By C. E. Taylor, L. V. Fegan

With on-stream electrical-resistance measurement as a basis, Bethlehem Steel developed a system for the completely automatic control of moisture in the feed to pelletizing. This system was later successfully adapted to the sintering process, with automatic control of moisture serving to maintain the air permeability required for effective burnthrough. Control of moisture content is an important factor in the agglomeration processes by which iron ores are converted into either pellets or sinter for use in blast furnace burdens. In both pelletizing and sintering the iron ore feed must be mixed with the proper amount of water so that capillary attraction and surface tension forces can be utilized to cause the ore particles to coalesce and form larger agglomerates. The success or failure of the entire pelletizing or sintering process depends on maintaining the proper moisture. Therefore, continuous automatic control of moisture is one of the conditions for the automation and optimum use of both pelletizing and sintering processes. In pelletizing, variation in moisture can cause ball size to change and is a common cause of cycling which results in a fluctuating feed to induration. Excessive moisture can cause lumps to form at the cutter bar and contributes to the formation of buildup on the green pellet screens. Green pellets with excess moisture deform and reduce bed permeability. In sintering, accurate control of moisture is also important because moisture must be within the limits that give the bed the required permeability for good burnthrough. In response to these problems, a moisture meter, based on the principle of electrical resistance measurement, was developed. This moisture meter has been a key factor in the development of systems for providing uniform feed of concentrates in most of Bethlehem's pelletizing and sintering plants. The automatic moisture control system for pelletizing was successfully plant-tested and resulted in the following: (1) Control of moisture in the green pellets to +0.2%. (2) Up to 50% reduction of bentonite consumption. (3) Control of unscreened green pellets so that 85% of the furnace product at Cornwall was within the size range of -5/8 +3/8 in. (4) Improved quality of product with a tumble test of 94 and a "Q" index of 90 as the result of feeding a consistently uniform feed to a shaft furnace. In sintering, where sinter-mix temperature change and variable feed could cause errors in moisture measurement, the system was designed to maintain the correct moisture under virtually all plant operating conditions. It was successfully plant-tested and is now in use in most of Bethlehem's sintering plants. The system achieves: (1) Control of moisture within the desired limits for optimum production. (2) Continuous measurement and control of moisture even with intermittent feeding of mix, so that attention by the operator is seldom required. The two automatic control systems provide smoother startup, minimize upsets and contribute to the consistent control of process variables in pelletizing and sintering. MOISTURE-CONTROL SYSTEM FOR PELLETIZING Fig. 1 shows the general arrangement of the system for automatic control of moisture in pelletizing.1 In this system a bin level controller, using a single probe which regulates filter speed, maintains a uniform low concentrate level in the bin.' The moisture meter, comprised of an analyzer and control components, continuously measures the moisture in the bed and adds water to adjust to the desired level. The meter is placed in an open-loop (feed-forward) control, i.e., water addition occurs downstream from the probes. Filter vacuum must be set so as to produce a filtercake moisture which is below the optimum for pelletizing. Bentonite is added after the

Jan 1, 1969

By H. R. Hose

The total crude fluorspar reserves in Western Europe, including the USSR and the Soviet sphere, Thetotalare estimated to exceed 17 million metric tons. The total fluorspar production in 1951 in Western Europe, excluding the USSR and Soviet sphere, amounted to 371,109 metric tons. The chief fluorspar-producing countries in Europe, excluding Russia and the Soviet sphere, are France, Western Germany, Italy, Spain, and the United Kingdom. TOTAL crude fluorspar reserves in western Europe, containing more than 35 pct CaF2, are estimated at 12 million metric tons, while reserves in the USSR and Soviet sphere may exceed 5 million metric tons. The bulk of these reserves are concentrated in the USSR, Spain, United Kingdom, Western Germany, and Italy. Since published estimates of the tonnages of individual mines and explored deposits are not available for most countries in Europe, the estimates given in Table I for individual countries are necessarily approximate, as they have been compiled from various sources and modified and combined according to the author's judgment. The majority of fluorspar deposits in Europe belong to the late Paleozoic metallogenetic epochs and have been found within or along the borders of the Hercynian massifs. Deposits in the Central and Esterel massifs of France, those of Western and Eastern Germany, Sardinia, the Catalonian coastal range and the Sierra de los Santos in Spain, and those of Devonshire and Cornwall belong to the Hercynian metallogenetic epochs. The deposits in southern Norway and Sweden belong to the middle Paleozoic or later, those of northern England to the late Paleozoic or later, those of northern Spain and Greece to the late Mesozoic, and those of northern and probably central Italy and Switzerland to the early Tertiary. Fluorspar occurs typically in veins, lenses, and stringers, but also in bedded replacement deposits and pockets, and may be colorless to white, yellow, green, pink, brown, blue, or purple to almost black. Except for the colorless variety, the only kind which can be used for the optical industry, the color of the fluorspar does not affect its ultimate use. Crystalline cubic or octahedral fluorite of CaF2 composition, with 48.9 pct fluorine and 51.1 pct calcium, has a specific gravity of 3.18. The ore in place, however, usually averages 3.0 metric tons per cubic meter owing to the quantity of other minerals asso- ciated with the fluorspar in the veins and the physical state of the fluorspar, which may vary from sugary to massive crystalline or microscopic inter-growths with barite or quartz. Barite of 4.5 sp gr, quartz of 2.66 sp gr, and chert and calcite are commonly associated with fluorspar. Other minerals usually present in lesser amounts are celestite, wad, silver, lead, zinc, copper, iron, and uranium minerals. In some cavity fillings the rhythmic deposition of the fluorspar, barite, and quartz into successive crusts may make it possible to handpick the crude ore on belts or to separate the fluorspar from other minerals while the actual mining is carried on. Where metallurgical grade lump spar is required for foundries, separation of quartz and barite may be made also by jigs and heavy media, and in some mines the lead and zinc sulphides may be separated and recovered by vibrating tables. However, except in the case of unusually high grade deposits, acid grade fluorspar is obtained by flotation. By this process fluorspar deposits grading down to 35 pct CaF2 in the crude ore can be worked at a profit when fluorspar, lead, zinc, and silver prices are high, as the sulphides of these ores can be separated from the fluorspar by flotation and concentrated.

Jan 1, 1956

By Gordon H. Geiger, John M. Svoboda

One mechanism of metal penetration into mold aggregates by cast steels is vapor state mass transport. In order to further understand and quantify this mechanism, the steady-state diffusion of metallic vapors was studied at 1250" to 1600°C in controlled oxygen pressure atmospheres by measuring the weight loss from a metal bath whose vapors were allowed to diffuse through a porous aggregate compact held at constant temperature. Mass transport rates substantially greater than those predicted from the assumption that the vaporizing species was the elemental metal vapor were observed. Vaporization rates were found to increase with an increase in the oxygen partial pressure in the gas atmosphere with maximum vaporization occurring at oxygen partial pressures just below those in equilibrium with a liquid oxide layer at the given temperature. When the liquid oxide layer was present, no vaporization was observed. Metal carbonyls were ruled out as the vapor species involved because of their thermodynamic properties. The species with the enhanced volatility were deduced to be gaseous metal-oxygen molecules, probably the polynuclear oxides (FeO)2 and (Mn0)2. Diffusion followed the ordinary diffusion mode, and Knudsen diffusion is not important in normal foundry molds. The interdiffusivities of the metallic vapors and the mold atmosphere, through the porous aggregates, were measured. The effective interdif-fusivity of the vapor species and mold atmosphere through the aggregate may be expressed as follows: Deff,12 - — where E is the void fraction of the aggregate and 7 is the "tortuosity factor". Aggregate material had no effect on diffusion rates, and the only relevant parameter was the void fraction. Values for 7 ranged from 3 to 6 for normal foundry aggregates with 4 being considered a good representative value. An area of interest in high-temperature metallurgical processes which has received little attention in the past is the diffusion of metal vapor species in porous aggregates. Research into the metal penetration phenomenon' in foundry molds has shown that a major penetration mechanism is the vapor transport of metallic species into the porous foundry sand molds, followed by condensation of metallic films on the sand grains. Considerable study of the vaporization of metallic species has been reported in the literature, and the vapor transport properties of many metallic species have been determined.2'3 Many instances have been observed where the vaporization behavior could only be explained by the existence of a metallic species with a vapor pressure greater than the elemental metallic species. These species were identified or, in some cases, presumed to be carbonyls,4,5 hydroxides,6-9 or vaporous oxides.10-16 The diffusion of gases through porous aggregates has received considerable attention in the literature pertaining to catalysis, although primarily at low temperature, and more recently in the literature concerned with ore reduction.17-22 The following experiments were devised and carried out in order to determine the nature of the metallic species involved in the metal penetration phenomenon, and the effective diffusivities involved. EXPERIMENTAL APPROACH Steady-state diffusion was selected for experimental convenience. The experimental systems used, shown in Figs. 1 and 2, consisted of a porous aggregate specimen contained in an impermeable alumina tube held above, and just separate from a metal bath. The entire system was held at constant temperature and pressure in a furnace with a controlled oxygen partial pressure atmosphere. After a brief period of nonsteady state behavior, a steady flux of metal vapor through the porous aggregate will develop due to the concentration gradient of metallic vapor in the gas phase. The flux of metallic vapor from pressure may be neglected because there is no pressure gradient. There is no flux from convection because there is no net flow of gas within the metal-aggregate system, and thermal diffusion is negligible compared to the flux from ordinary diffusion. If the material being studied is allowed to vaporize until steady-state behavior is attained, the experimental results may be analyzed using Fick's First Law.

Jan 1, 1970

By D. E. Baldwin

A method has been devetoped for predicting the produced concentration profile for a miscible slug in a five-spot pattern. The technique consists of dividing the five-spot into radial etements and applying an approximate radial solution of the dispersion equation to each element. To test the method, laboratory experiments were conducted in five-spot models. The models were 2-ft X 2-ft X 2-in. Berea sandstone slabs simulating one quarter of a five-spot. An aqueous system was used with tritiated water as the tracer fluid. Effiuent concentration profiles were obtained for slug sizes ranging from 0.7 to 10.2 per cent of the pore volume. Agreement between predicted and experimental profiles was excellent. A prediction was also made for a field tracer test. For this case the predictive technique was modified to account for stratification by use of a layer cake model. INTRODUCTION Dispersion in porous media is of growing interest to the petroleum industry because of the increasing importance of secondary recovery operations. One of the controlling factors in the recovery of oil by miscible displacement or by the use of waterflood additives is the degree of mixing between the fluids of interest. Consequently, one of the prerequisites of any recovery prediction is an adequate method of predicting dispersion. Also, any quantitative interpretation of waterflood tracer tests is dependent upon the ability to determine the dispersion of the tracer as it flows through the reservoir. During the past few years there have been reported in the literature the results of a large number of investigations, both theoretical and experimental, concerning dispersion in porous media. (An excellent review of these results has been presented by Perkins and Johnston." Only a few of these investigations, however, deal with non-lin-ear flow systems. Raimondi et al.9 have presented an approximate solution of the dispersion equation for radial, diverging flow. Lau et al. 7 and Bentsen 1 have presented laboratory data for radial systems and have shown that the dispersion can be adequately predicted by Raimondi's solution. Bentsen has also presented a small amount of data for a radial, converging system. Brigham and Smith' have applied Raimondi's solution, with approximating assumptions, to a five-spot pattern to predict the behavior of a field tracer test. The afore-mentioned investigations have been concerned almost entirely with radial diverging flow and have neglected the effect of converging flow. For this reason the present investigation was undertaken to develop a method of predicting dispersion in a five-spot pattern including the effects of both diverging and converging flow. The method which was developed consisted of dividing the five-spot into diverging-converging radial elements and applying an approximate solution of the disperson equation to each element. DEVELOPMENT OF PREDICTION METHOD The development of the prediction method consisted of two phases: construction of a radial model of a five-spot pattern and derivation of equations to describe the dispersion in the radial model. FIVE-SPOT MODEL A radial model which simulated the flow characteristics of a five-spot pattern was constructed by dividing the five-spot into diverging-converging radial elements (Fig. 1). Because of symmetry the model represented only an octant of a complete five-spot. Each element had an included angle of l° (making a total of 45 elements) and a radius r, such that the sum of the individual element areas was equal to the area of the five-spot octant. The normal five-spot fractional flow behavior was determined by averaging the experimental five-spot data of Caudle and Witte,' Dyes, Caudle and Erickson 5 and Fay and Prats. 4 The resulting curve is shown as a solid line in Fig. 2. To match this curve with the radial model it was

Jan 1, 1967

By Robert E. Green, Edmund G. Henneke

Expressions are derived for the orientation factors associated with dislocation damping due to dislocation motion in the basal slip plane. The results of the calculation of these factors for zinc crystals aye pvesepzted and the relevance of these results to ultvasonic-attenuation measuvements is discussed. In this paper expressions are derived for the orientation factors associated with damping of sound waves due to dislocation motion in the basal slip plane of hexagonal crystals. For zinc crystals, the orientation factors for dislocation motion in one slip direction only, as well as for simultaneous motion in all three basal slip directions, have been calculated at 1-deg intervals. The results of these calculations are presented in the form of contour diagrams in a sector of a standard polar coordinate graph. The calculations have been carried out for both the traveling- and the standing-wave conditions. The relevance of these results to ultra sonic-attenuation measurements is discussed. Associated with the calculation of the orientation factors, the modes of elastic-wave calculation were also evaluated, according to the procedure previously employed by Hinton and Green.' The results of these calculations are not presented here since similar results have been reported earlier by Mus-grave.2 However, it should be mentioned that, although the qualitative features of Musgrave's results are preserved, the quantitative values are somewhat altered since different values for the elastic constants of zinc were used. ORIENTATION FACTORS It is convenient to refer to three different sets of coordinate axes, Fig. 1, in the following calculations. First, there are the Miller-Bravais axes which are used in describing the Miller-Bravais indices of planes and directions in a hexagonal crystal. These axes are denoted in the conventional manner by a,, a,, a,, and c. The second set of axes used is a rectangular Cartesian coordinate system located as shown in Fig. 1. Note that the x axis coincides with the close-packed direction al [2110], the z axis coincides with the c axis [0001], and the y axis is orthogonal to these two lying in the [0110 ] direction. Finally, a form of Eulerian angles is used to locate the position of the specimen axis. The angle between the specimen axis and the c axis is denoted by 0 and the angle between the projection of the specimen axis on the basal plane and the a, direction is denoted by ø. The majority of observations on plastic deformation of metal crystals indicates that such deformation is caused by the motion of dislocations on a limited number of crystalline planes and in a close-packed direction. In crystals with hexagonal symmetry the most prominent slip plane is the basal plane (0001) and the slip directiops- are one of the close-packed directions, [2110], [ 12101, or [ 1120]. To properly set the stage for the following work, several points must be clarified. First, we want to consider the contribution to the attenuation of elastic waves by dislocations only. While the Granato-Lucke, model for dislocation damping is presently the one most quoted, it is not necessary to refer the present work to it. The only assumption that need be made is that the motion of dislocations absorbs energy from the elastic wave and therefore attenuates it. It is immaterial here how this energy absorption occurs. However, because dislocations have been observed to move most freely on the (0001) plane for hexagonal crystals undergoing plastic deformation, it will be assumed that it is the dislocation motion on this plane which attributes to the observed attenuation. The basic orientation factor, Ok, is essentially the same as that used by Green and Hinton for fee crystals,4 defined by the equation

Jan 1, 1968