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By P. M. Blair, D. W. Peaceman

The shape and position of the gas-oil transition zone during downdip displacement of oil by gas has been calculated using flow equations which include the effects of gravity, relative permeability, capillary pressure and compressibility of the fluids. The calculations treat the problem in two space dimensions, and results are compared with data from a laboratory model tilted at 30° and 60° from the horizontal on displacements near and above the maximum rate at which gravity segregation prevents channeling of the gas along the top of the stratum. The good agreement between calculated and experimental results demonstrates the validity of the tecbnique as well as that of the flow equations. INTRODUCTION Knowledge of the fluid distribution and movement in an oil reservoir is important in producing operations and estimation of reserves. The history of the oil industry has included steady progress in improving the accuracy of calculations which provide the required knowledge. The earliest method of calculating reservoir performance consisted of material-balance equations based on the assumption that all properties were uniform throughout a reservoir.l For many reservoirs such a simple formulation is still the most useful. However, when large pressure and saturation gradients exist in a reservoir, the assumption of uniform values throughout may lead to significant error. To reduce these errors, Buckley and Leverett2 introduced a displacement eqation which considers pressure and saturation gradients. Methods available at that time permitted solutions to the Buckley-Leverett equation in one space dimension; these solutions have been very useful in solving many problems related to the production of oil. However, the one-dimensional methods are not adequate for systems in which saturations vary in directions other than the direction of flow. An example of such a system is the case of gas displacing oil down a dipping stratum in which the gas-oil contact becomes significantly tilted. Of course, the Buckley-Leverett displacement method cannot predict the tilt of the gas-oil contact. Recent improvements3,4 of the one-dimensional Buckley-Leverett method achieve some success in predicting the tilt of the gas-oil contact at sufficiently low flow rates. However, at rates high enough that the viscous pressye gradient nearly equals or exceeds the gravity gradient, even these improved one-dimen,sional methods incorrectly predict the shape and velocity of the contact. Further progress in estimating such fluid movements in a reservoir appears to require consideration of the problem in more than one space dimension. The recent two-dimensional method of Douglas, Peaceman and Rachford5 appears adaptable to calculate changes with time of the saturation distribution in a vertical cross-section of a reservoir. The movement of saturation contours should represent the moving fluid contacts and include the effects of crossflow due to gravity, as well as variations in the rock and fluid properties. The nonlinear nature of the equations used in the method has prevented proof of the validity of the solutions. Douglas, Peaceman and Rachford made some comparisons with experiment but did not include cases in which gravity was important nor cases involving displacement by the nonwetting phase.For these reasons, a test of the two-dimensional method for a case in which these factors are included would be very desirable. The test selected was a comparison of calculated results with those from a carefully controlled laboratory experiment on a model with measured physical properties. The model selected was one in which gas displaced oil down a tilted, rectangular sand pack. The model can be thought of as representing a vertical cross-section taken parallel to the dip of a reservoir. The displacement thus simulates gas displacing oil downdip that might result from gas-cap expansion or gas injection. This report outlines the calculation method used

By R. D. Saltsman

This paper deals with the evolution of a test procedure devised to totalize a coal's ignitibility, its degree of caking, and its coke reactivity and the development of a method for assigning proper combustion indices to stoker coals. The application of this type of work to stoker design is also discussed. It is believed that this work may be the prelude to understanding the effect of various petrographic constituents on the combustion characteristics of a coal or a blend of coals. The inadequacy of our knowledge of the coal combustion process has long been recognized. This was expressed very well by Spicer and Finney.1 "The control of combustion processes in industry and the design of the equipment necessary for efficient burning to take place have always been very much branches of an industrial art rather than of an exact science. This has been especially true of the utilization of coal, a substance so heterogeneous and complex in character that all too little is known about its chemical and physical composition." During the past several years one branch of coal utilization, that of high temperature carbonization, has made significant advances from being an art into being an exact science. This has been accomplished by the development of petrographic techniques for identifying, measuring, and determining the thermal properties of the organic materials which affect the carbonization of coal and the correlation of this information with a pre-existing standard of coke quality, the stability index. The success in using petrographic analysis for predicting the coking properties of coals and their blends suggested that petrographic techniques might be applied to the combustion process for predicting the combustion characteristics of coals. Certainly the different organic entities2 influence the combustion characteristics of a coal as they influence its coking properties. If it were possible to predict combustion characteristics from an analysis of a coal's petrographic composition, it would seem logical that an intelligent investigation could then be made into methods of altering stoker coals to improve their combustion characteristics. The successful application of petrographic analysis in the field of high temperature carbonization was made possible by the pre-existence of the stability index, which is a numerical value of coke quality. The quality of combustion is determined by the sum of the effects of the ignitibility, the caking, and the coke reactivity characteristics of a coal. While the literature is filled with papers reporting the correlation of various analyses to the ignitibility, or to the caking, or to the coke reactivity characteristics of coals, none of these earlier studies could be related to one index. The lack of a single index for combustion quality precluded the application of petrographic techniques for predicting combustion characteristics. SELECTION OF THE COMBUSTION TEST EQUIPMENT The primary prerequisite of any test developed to evaluate the relative performance of stoker coals must reflect the three characteristics of the coal burning process: 1) ignitibility, 2) degree of caking, and 3) coke reactivity. The widespread variations in these combustion characteristics of the available stoker coals are well known. The ignitibility, which has been defined as that characteristic which determines the ease with which the fuel may be brought to a condition of self-supporting active oxidation, varies from where a coal is deemed to be flashy to the opposite extreme of being fireproof. The ignitibility of a coal influences its ability to pick-up from hold fire, greatly affects the degree of bed stability during swing loads, and is a factor in determining the average burning rate obtainable on a given grate. The caking characteristic affects the degree of bed porosity during the devolatilization period, and coals are classified on a scale varying from free burning to highly caking. Caking is not synonymous with bed swelling since a bed may swell two to three times its normal depth during the period of devolatilization yet remain very porous, allowing an efficient and high rate of combustion to take place. Caking is melting

Jan 1, 1964

By J. N. Ong

The oxidation of tantalum is assumed to occur by four simultaneous first-order chain reactions; solution of oxygen in the metal, nucleation and growth of a suboxide phase at the metal surface and two phase boundary processes which give rise to two different modifications of Ta2O5. A rate equation is then formulated which accounts for all principal details of oxidation behavior, unifies and correlatss the results of the principal investigations, and expresses the rate as a function of pressure, temperature, and time. For purposes of practical application, the expressions: may be used below and above 700T respectively to determine the regression rate of tantalum in cm per hr, where T is to be expressed in "K and P in atmospheres pressure of oxygm. The equations will reproduce within a factor of two all reported experimental rates in the temperature range 475" to 1400°C and pressure range 0.000026 to 40.8 atm 0. 1 HE oxidation kinetics of tantalum have been investigated thoroughly in the temperature range 50" to 1400 and pressure range 1.32 x l0-' to 40.8 atm pressure of 0,and as a result, the principal features of the oxidation of tantalum are known and have been described in considerable detail. The exposition of the characteristics of the behavior of tantalum in oxygen has been largely descriptive, with considerable attention being devoted to elucidating the structure, composition, and morphology of the resulting oxide products. The weight change vs time results have usually been reported graphically. With the exception of the linear rate below 700 C and the process of diffusion of interstitial oxygen into the metal,6'9 no effort has been made to characterize the reaction rates in functional form. It is the purpose of this paper to demonstrate that when the fundamental principles of chemical kinetics are rigorously followed, the results of the principal investigations will be correlated and unified and the oxidation rale may be expressed as an analytic function of pressure, temperature, and time. In addition to accounting for the temperature and pressure dependency of the oxidation rate, the following minor features will be accounted for as a result of the basic analysis: the initial period of nonline-arity in rate at low temperatures, which decreases with increasing temperature;3'5'7 the rate vs temperature cusps obtained at constant pressure"5 the concentration gradient of oxygen interstitially dissolved in the metal.5'7 The analysis will also be consistent with the overall chemistry of reaction and with all visual, microscopic, and X-ray observations. DEVELOPMENT OF RATE EQUATION As a fundamental basis for the treatment of metal-oxygen reaction kinetics, it will be regarded that the rate of reaction is determined by the concentration and not the activity or chemical potential. It is further assumed that all metal-oxygen reactions are first order complex chain reactions" whose rate is dependent on the concentration of an intermediate species which is a complex function of the oxygen concentration, or oxygen pressure, only. The basic rate equation is therefore: is the weight of oxygen consumed by the reaction, A is that portion of the area of the metal over which the reaction is occurring (cm2), t the time (sec), kf the rate constant for forward reaction, C the intermediate species concentration and f '([Oz]) and f (Po,) are complex functions of oxygen concentration and Oxygen pressure respectively. Reverse reactions may usually be neglected in metal-oxygen systems since a large negative free energy generally accompanies these reactions. Eq. [I] takes different forms when the rate of reaction is controlled by diffusion into or through a phase or at a phase boundary. For diffusion into the metal phase, the rate is given by1' Di is the diffusion coefficient of species diffusing into the metal phase (cm2/sec), and Q is the corre-

Jan 1, 1962

By Arthur H. Wilson

INTRODUCTION The stratified deposits of the coal measures are strongly laminated and exhibit a high degree of anisotropy. The stronger rocks in the sequence possess joints and shear planes, the coal itself has a closely spaced cleat pattern, and the softer rocks are capable of flowage. To calculate the general distribution of stress in such a medium by analytical means based on the theories of elasticity is difficult, and has not yet been satisfactorily achieved. Nevertheless, some estimate is required to aid the design of supporting pillars and permit the correct siting of main roadways. This is important when laying out districts for longwall extraction, particularly if the results of previous experience are not available in the area to be exploited. Although elastic stress analyses, including finite element analysis, may not be very appropriate, the basic equilibrium conditions will apply. These, combined with logical deductions based on observed fact, can be used to give appropriate solutions which experience has shown to be capable of giving acceptable results in the design of pillars associated with longwall workings. Constancy of cover load is one of the fundamental equilibrium conditions. When mining below ground, the cover is not removed, therefore the average vertical load after mining must equal the initial cover load. If mining produces a local decrease in load in any particular area, it must be balanced by an equivalent increase in load nearby. Bending moments in the strata above are capable of compensating for the turning moments which this redistribution of load may produce, and the existence of a "rear abutment", much discussed between 1950 and 1960, is not an essential equilibrium requirement. If the distribution of vertical stress is plotted three-dimensionally, then the load on any small element of area 6S will be o.6S, where o is the average stress. This also represents the volume of the "stress envelope" which lies above 6s. By making the individual areas very small and summing together, it is possible to show that the total load on any particular area can be found by calculating the volume of the stress distribution envelope which lies above it. A similar consideration applies to the stress distribution in a two-dimensional situation. If the stress distribution curve is drawn, then the area below any portion will represent the load per unit depth of section. For example, Figure 1 represents the possible stress distribution across a caved waste and associated ribside. If AW is the load deficiency [ ] These loads must balance, therefore area AW = AS, and if one can be deduced, the other will then be known. Attempts have been made to measure stress distributions in the ribside and in the waste. Stress meters in soft rocks such as the coal measures have not proved very effective. Those based on the measurement of strain have experienced difficulty in strata capable of flowage, high modulus stress meters suffer from cross sensitivity, and overcoming to obtain absolute stress is almost impossible. Load measuring instruments left in the waste have also proved largely ineffective, mainly as a result of the difficulties of protecting the leads in a caving environment. Certain logical deductions, however, can be made about the conditions in the waste. CONDITIONS IN THE WASTE Over the area of the waste the surface trough which develops is remarkably consistent, despite the geological variations which occur from coalfield to coalfield. In Britain, an accuracy of 210% is claimed for subsidence prediction based on width of extraction and depth of cover (National Coal Board, 1975). The subsidence in the USA is more variable (Adamek and Jeran, 1981), particularly with respect to the point of inflexion, but the same general pattern occurs. At the shallower depths the somewhat harder roof strata involved in the caving in the USA may suffer less compaction, giving reduced values of subsidence. Subsidence at intermediate horizons below the surface

Jan 1, 1982

By M. M. Fine

THE problem of producing a uniform, medium-fine sand for glass-furnace feed has been of interest to the glass-container industry for many years. In the present investigation of the problem, conducted by the Bureau of Mines in cooperation with the Owens-Illinois Glass Co., Alton, Ill., a satisfactory method of producing such sand was developed. The investigation involved continuous-grinding studies of glass sand with flint and porcelain pebbles, steel balls and rods, and a simultaneous study of classification. The most satisfactory procedure consisted of grinding in a pebble or ball mill followed by a double classification: hydraulic classification to separate the oversize for return to the grinding unit and mechanical classification to remove fines and impurities from the ground sand. The proper application of this flowsheet usually yielded a graded sand containing less iron than the feed even when grinding was done in a steel-lined ball mill. The amount of sand ground into the so-called "dusting" sizes was, in all cases, less than 10 pct. Experiments were made which showed that this very fine sand could be beneficiated to a usable material. Background The advantages in utilizing a uniform, medium-fine sand as a feed for glass furnaces has motivated a considerable amount of research on the ,problem of producing such sand by glass producers and others. For the glass producer, the faster melting rate of such a graded sand offers the possibilities of reduced fuel costs and increased furnace capacity. In addition, there is the likelihood of improving the quality of the melt and thereby reducing the number of seeds (air bubbles) and stones (unrnelted grains) in the finished ware. Although the melting rate of a glass batch depends on other factors in addition to the grain size of the sand, there is experimental evidence to show that grain size influences the melting rate considerably. Potts, Brookover, and Burch' found that when the raw materials (sand, soda ash, and limestone) were always matched in- size, there was a great decrease in the melting time as the grain size of the materials was decreased. They found, too, that when the grain sizes of the soda ash and Limestone were held constant at 40 to 60 mesh, the minimum melting time occurred when 40 to 60 mesh sand was used. The increase in melting time with decrease in grain size beyond 40 to 60 mesh was due, they postulated, to accelerated demixing or separation of batch during melting as the grain size of the sand decreased in relation to the grain size of the other materials. An earlier investigatora found that the highest rate of melting and purifying ordinary glass was obtained when medium-size sand (0.385 to 0.120 mm, correspondingly roughly to 40 to 120 mesh) was used. Although batches containing smaller-size sand were found to fuse more rapidly, the time required to remove air bubbles from such melts actually increased the total duration of the process. Coarsegrained sand, it was stated, increased greatly the times of both melting and purifying. Although the laboratory data are difficult to confirm on a plant scale, observations by the Owens-Illinois Glass Co. indicate a higher furnace capacity of better-quality ware when a medium-fine sand is used in the batch. During the recent war, the glass producer was inclined to overlook a considerable variation in raw materials in the interest of higher production. For instance, table I shows the wide range in the amounts of various screen fractions in sand supplied by one producer to the Terre Haute plant of Owens-Illinois. The sand, from the northern Illinois St. Peter sandstone, represented material used by that plant over a period of a few years during the war. Since the war, the glass industry has insisted on

Jan 1, 1951

By M. M. Fine

THE problem of producing a uniform, medium-fine sand for glass-furnace feed has been of interest to the glass-container industry for many years. In the present investigation of the problem, conducted by the Bureau of Mines in cooperation with the Owens-Illinois Glass Co., Alton, Ill., a satisfactory method of producing such sand was developed. The investigation involved continuous-grinding studies of glass sand with flint and porcelain pebbles, steel balls and rods, and a simultaneous study of classification. The most satisfactory procedure consisted of grinding in a pebble or ball mill followed by a double classification: hydraulic classification to separate the oversize for return to the grinding unit and mechanical classification to remove fines and impurities from the ground sand. The proper application of this flowsheet usually yielded a graded sand containing less iron than the feed even when grinding was done in a steel-lined ball mill. The amount of sand ground into the so-called "dusting" sizes was, in all cases, less than 10 pct. Experiments were made which showed that this very fine sand could be beneficiated to a usable material. Background The advantages in utilizing a uniform, medium-fine sand as a feed for glass furnaces has motivated a considerable amount of research on the ,problem of producing such sand by glass producers and others. For the glass producer, the faster melting rate of such a graded sand offers the possibilities of reduced fuel costs and increased furnace capacity. In addition, there is the likelihood of improving the quality of the melt and thereby reducing the number of seeds (air bubbles) and stones (unrnelted grains) in the finished ware. Although the melting rate of a glass batch depends on other factors in addition to the grain size of the sand, there is experimental evidence to show that grain size influences the melting rate considerably. Potts, Brookover, and Burch' found that when the raw materials (sand, soda ash, and limestone) were always matched in- size, there was a great decrease in the melting time as the grain size of the materials was decreased. They found, too, that when the grain sizes of the soda ash and Limestone were held constant at 40 to 60 mesh, the minimum melting time occurred when 40 to 60 mesh sand was used. The increase in melting time with decrease in grain size beyond 40 to 60 mesh was due, they postulated, to accelerated demixing or separation of batch during melting as the grain size of the sand decreased in relation to the grain size of the other materials. An earlier investigatora found that the highest rate of melting and purifying ordinary glass was obtained when medium-size sand (0.385 to 0.120 mm, correspondingly roughly to 40 to 120 mesh) was used. Although batches containing smaller-size sand were found to fuse more rapidly, the time required to remove air bubbles from such melts actually increased the total duration of the process. Coarsegrained sand, it was stated, increased greatly the times of both melting and purifying. Although the laboratory data are difficult to confirm on a plant scale, observations by the Owens-Illinois Glass Co. indicate a higher furnace capacity of better-quality ware when a medium-fine sand is used in the batch. During the recent war, the glass producer was inclined to overlook a considerable variation in raw materials in the interest of higher production. For instance, table I shows the wide range in the amounts of various screen fractions in sand supplied by one producer to the Terre Haute plant of Owens-Illinois. The sand, from the northern Illinois St. Peter sandstone, represented material used by that plant over a period of a few years during the war. Since the war, the glass industry has insisted on

Jan 1, 1951

By Blake M. Caldwell, Harrold D. Owen

The continued use of permanent-type, we11 completion Iras pointed LIP the need for more powerful through-tubing perforating equipment. A new expendable shaped charge perforator has been developed in which the charges are run through the tubing in a vertical position and then opened out to a horizontal position when at the desired shooting zone. The performance of this new tool is comparable to that of conventional casing type perforators. This paper describes the new tool and its applications. Performance data in targets rind under various actual well conditions are presented. INTRODUCTION The general acceptance and increasing use of pernianent-type well completion by the oil industry has indicated the need for through-tubing perforating equipment with performance comparable to that of conventional casing type perforators.'.' To achieve the performance desired, a new, expendable, shaped charge perforator has been developed with adequate power for effectively perforating both 51/2-in. and 7-in. casing below tubing. This has been accomplished without resulting severe damage to the casing and can he obtained under extremes of pressure and temperature. The design of this new tool is a radical departure from previous concepts of tubing guns since it was felt that the two major limitations imposed by previous designs severely handicapped performance. The first of these limitations was the restriction imposed upon the length of the charges by the necessary small inside diameter of the carrier. This resulted from the necessity for compromises in designing a hermetically sealed gun small enough to allow free passage through 2-in. tubing, yet with sufficient wall thickness to withstand some minimum pressure. In an effort to overcome this limitation, the charges were in some cases placed at an angle with the axis of the well with resulting loss in effective depth of penetration. The second limitation concerned the extreme distance of the charges from the casing when the gun was fired. An example of the latter situation is demonstrated by the fact that a conventional 13/4-in. tubing gun centralized in 51/2-in., 17-lb casing has a clearance of more than 1-9/16-in. on all sides. This means that a considerable portion of the force from each charge must be expended in penetrating this 1-9/16-in. of well fluid before it even reaches the casing. With the gun centered in 7-in., 23-lb casing, the amount of fluid penetration required of each jet is in excess of 21/4-in. In a situation where the 13/4-in. gun is touching one side of the 51/2-in., 17-lb casing, the jets from the charges on the opposite side have more than 3%-in. of fluid to penetrate. A similar situation in 7-in., 23-lb casing necessitates the penetration of more than 4-9/16-in. of fluid. The penetrating effect of the charges is severely impaired in passing through so much fluid. Under such conditions, there is usually insufficient force left to penetrate the casing. Previous studies have shown the decided disadvantage of having to shoot through such an excessive quantity of well fluid.'.' The problem. then, was how to position large powerful charges close to the casing prior to firing and still manage free passage through 2-in, tubing. The solution appeared to be in some method of folding sealed charge units in an open type, expendable. 13/4-in. carrier for travel through the tubing, and unfolding them upon their arrival at the shooting zone in the casing. It was desired to control the mechanism electrically from the surface through the regular shooting cable, making the release of the charges independent of the physical characteristics of the well, yet with a provision to prevent the gun from firing until the

Jan 1, 1955

By J. Szekely, V. Stanek

Steady state heat transfer and fluid flow rates are predicted for the counter-current contacting of gas and solid streams. For a .fixed pressure drop across the bed the convective heat transfer rate is affected by the gas flow rate, which in turn is affected by the temperature distribution in the system. Variations in bed voidage and in particle size were found to change the predicted temperature profiles quite markedly, in particular when the thermal capacities of the gas and solid streams were equal. IN Part I we examined the effect of step and pulse changes in the inlet gas temperature on transients in the flow of gases through fixed packed beds. Here we shall discuss heat flow and fluid flow interactions in the counter-current heat exchange between packed solids and a gas stream. Counter-current heat exchange in packed beds has been extensively studied. In his now classical paper, Furnas' assumed constant property values and developed a simple analytical. expression for the temperature profiles in the gas and in the solid stream. A highly sophisticated analysis of counterflow systems, allowing for transients and interparticle conduction was developed by Amundsen. In preparing a very readable review of this topic Elliott' also presented a number of elegant solutions, including the case involving heat generation. A great deal of work has been done on the application of these concepts to the modelling of the blast furnace. However, in all this previous work the mass flow of the gas was regarded as a quantity independent of the gas-solid heat exchange process. In the following we shall consider counter-current heat exchange between a gas stream and a packed bed of solids, under conditions where the pressure drop across the bed is specified, rather than the mass flow of the gas. FORMULATION Consider a packed bed, extending from y' = 0 to y' = L, in which moving solids and a moving gas are contacted in a counterflow arrangement. Let cold solids, with an initial temperature Tsi and mass flow GA, enter at y' = L, and hot gas with an initial temperature of Tgi and mass flow Gg enter at y' = 0. In contrast to the problems considered in Part I this system has a nontrivial steady state and we shall restrict our attention to the study of the pressure drop flow relationships and the resultant temperature distributions for steady state conditions. Under these conditions the formulation of the problem requires the statement of the following: 1) heat balance on the gas 2) heat balance on the solids 3) the mechanical energy balance equation and in addition the knowledge of the equation of state, and a functional relationship between viscosity and temperature and an expression for the heat transfer coefficient in terms of the other parameters. On neglecting temperature gradients within the solids and for a constant mass flow of the gas, the solution to this problem is well known.'" The expressions for the temperature profiles of the gas and solid streams may be put in the following form: where Tg and T, denote the dimensionless gas and solid temperatures respectively, y is a dimensionless distance, and R is the ratio of the thermal masses. These quantities are defined as follows: For the particular case of equal thermal mass capacities by the gas and the solid, that is R = 1, the Eqs. [1]and[2] simplify to Eqs. (71 and[8], respectively: i.e., the resultant profiles are linear. As for most practical cases , solutions given by Eqs. [7] and [8] will contain only the inlet temperatures as parameters. For a fixed, predetermined value of C;b the above

Jan 1, 1970

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

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

Jan 1, 1958

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

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

Jan 1, 1958

By W. Mitchell, B. H. Bergstorm, C. L. Sollenberger

A unique compression testing machine was constructed to load individual 1/8 to I-in. spheres of glass, etc., at rates from 100 to 100,000 lb per min. During loading the applied load was continuously plotted us the deformation of the sphere-platen system. The load-deformation curve conforms to the theory of elasticity. The spheres were surrounded by steel retaining rings or by gelatin (to prevent secondary fracture). The product size modulus was determined from screen analyses. The data shows: 1) Energy input per unit mass is inversely proportional to the product size modulus (or directly proportional to the surface area per unit mass). 2) Increased loading rates slightly increase the energy input per unit mass required for fracture without affecting the preceding relationship. Theory and experiments show that much of the input strain energy is transformed into kinetic energy of fragments at fracture, this being a reason efficiencies based on "surface energies" are exceedingly small. This large kinetic energy component is usually dissipated as heat but may be partially reclaimed because impact of fragments against the surface of steel retaining rings causes additional comminution. There are currently several theories of crushing, all of which can be derived from the general energy equation of Gillilandl as was pointed out by Mitchellet al in 1954'and more recently by Rose and Sullivan. If the exponent of the equation dE-x-"dx is assigned values of 2,1, or 3/2, one obtains by intergration between suitable limits, the equations of Rittinger, kick,' and Bond,' respectively. Other investigators prefer to let the exponent be a variable parameter. To help resolve the controversy, the Allis-Chal-mers Research Laboratories set up a series of carefully controlled experiments in which the level of energy input required to induce the comminution of a single particle by slow compression loading was very accurately measured. The energy input requirement, in conjunction with the size analysis of the fragments produced at fracture, yielded an accurate energy-product size relationship for the conditions studied. APPARATUS The experimental apparatus was designed to apply a compressive load on opposite sides of an "ideal particle." Spherical shapes were chosen, because point contact loading is typical of the mechanics of conventional crushers. The bulk of the experimental data was obtained with glass spheres because glass is a fairly homogeneous, brittle material for which the properties follow definite statistical distributions. The spheres ranged from 1/8 to 1-in. diam. The compressive load was applied to the specimen at a uniform rate by a loading mechanism (Fig. 1) which acted through a system of levers. The uniform rates, ranging from 100 to 100,000 lb per min, were obtained by metering water into a bucket suspended at the end of the secondary level. Point contact was made with the sphere through tungsten-carbide platens. This latter material is extremely hard and has a high modulus of elasticity, thereby having a minimum of distortion when loaded. The actual load on the sphere was measured independently of the lever system by a load cell, Fig. 2, which was designed so that the voltage output of the load cell would be proportional to load. The deformation of the specimen was indirectly measured by a pair of transformers which were attached to opposite edges of the upper platen. The cores of the transformers, which were attached to the lower platen, moved relative to the transformer so that the voltage output was proportional to the deformation of the sphere-platen system. The load was plotted against the deformation on an X-Y recorder. The area under the curve, as shown on Fig. 3, was proportional to the energy input of the

Jan 1, 1961

By Henry A. Wriedt, John Chipman

Equilibrium in the reaction of hydrogen gas with oxygen in liquid nickel, iron, and their alloys has been studied at temperatures of 1500° to 1700°C. The equilibrium con^stant, 0/p, [% O], is greater in nickel then in iron by a factor of 45 at 1594 C. In the alloy steel range, the activity coefficient of oxygen is slightly increased by the presence of nickel. NICKEL is more resistant to oxidation than is iron, in both the solid and the liquid state. The solubility of oxygen,' however, is greater in nickel than in iron. The two metals and their liquid alloys constitute an interesting series of solvents in which the chemical behavior of oxygen can be studied by methods similar to those used for liquid iron.' Experimental Method The furnace assembly, including the tube for preheating the entrant gas, was the same as that described by Gokcen and Chipman. " controlled mixture of water vapor, hydrogen, and argon preheated to approximately the melt temperature impinged upon the surface of the liquid metal. The charge, weighing about 35 g, was held in an alumina crucible at constant temperature in the gas stream for a prolonged period, then quenched in a stream of cold helium, sampled, and analyzed for oxygen by the vacuum fusion method. The temperature of the liquid metal was measured by a Leeds and Northrup optical pyrometer which had been checked against a similar instrument certified by the National Bureau of Standards. The freezing point of electrolytic iron in hydrogen, taken as 153j°C, was used for frequent checks. This point and the emissivity determined by Dastur and Gokcen" established the transmissivity of the optical system. The emissivities of Fe-Ni alloys at 1535" are known from the experiments of Smith and Chipman,h nd it was assumed that the variation with temperature is parallel with that of iron. Recorded temperatures are probably accurate within k5"C. Preparation of Gas Mixtures—Since the ratio H,O:H, had to be much higher for nickel than for iron, it was found convenient to construct two separate systems for gas preparation. In both cases, the ratio of argon to the sum of the other two gases was at least 4, the proportion of hydrogen and water vapor being varied to produce H,O:H, ratios from 0.2 to 70. The primary purpose of the argon was to diminish the errors caused by thermal separation' in the gas phase. In addition, its presence lowered the saturation temperature required for a given H,O:H, ratio and minimized the difficulty of preventing condensation. Both systems were modifications of the one used by Dastur and Chipman.' For H,O:HI ratios greater than 2, welding-grade argon purified by passage over magnesium chips at 620°C was mixed with hydrogen produced in the electrolytic cell shown in Fig. 1. The design was based on a cell used by Bodenstein and Pohl," the important features of which area nickel wire cathode and sodium hydroxide solution as the electrolyte. The rate of hydrogen evolution was controlled by a calibrated ammeter and a rheostat in the circuit, the current efficiency being taken as 100 pct. Precise control of the argon flow rate at about 300 ml per min was achieved by means of a capillary flowmeter having a controlled pressure drop. The molar flow rate was measured volumetrically before each run, and a constant flowmeter setting was maintained thereafter. The A-H, mixture passed through a water saturator of the type previously described," the saturation temperature being maintained automatically with a fluctuation of k0.04"C. The tube between saturator and furnace was heated by a resistance winding. The system was of all Pyrex construction with the exception of the stainless steel tube containing magnesium chips. This was connected with the glass tubing at the upstream end by

Jan 1, 1957

By C. M. Wayman, R. Bullough

The selection rules fir twinning in fcc, bcc, and hcp lattices as established in Part I are used to predict the positions, relative to the matrix (untwinned) reciprocal lattice, of reflections due to double diffraction from the twins. These doubly diffracted reciprocal lattice points, which often coincide with ordinary matrix reciprocal lattice points, are deduced from kinematical considerations. 1. INTRODUCTION. A Bragg diffracted beam under certain conditions may be rediffracted to produce a third beam i .e., the incident beam, the first diffracted beam, and the rediffracted beam). In such a case an initial incident beam is diffracted by a set of planes; the diffracted beam may then be rediffracted by a second set of crystallographically different planes. The initially diffracted beam acts as an incident beam for the second set of planes, thus resulting in a twice-diffracted beam. This twice-diffracted beam may be regarded as a singly diffracted beam from a set of lattice planes whose structure factor is zero. In order to find the reciprocal lattice point corresponding to the doubly diffracted beam, it is customary1,2 to combine vectorially the component reciprocal lattice vectors which head to the extra reflections. For example, hcp metals may exhibit the forbidden (0001) reflections.* As Fig. 1 shows, the (0001) reciprocal lattice point results from: (1010) + (1011) = (0001) Similarly, combination of the beams (1011) and (1010) gives rise to: (1011)- (1010) = (0001) As Pashley and stowell2 have pointed out, even if in fact the double diffraction points lie on the Ewald sphere, they may not actually represent observed doubly diffracted beams because the initially diffracted primary beam, i.e., the source, must also correspond to a reciprocal lattice point which lies on the sphere of reflection; however, the reciprocal lattice point corresponding to the planes causing the rediffraction need not lie on the sphere. A little consideration of the Ewald construction in Fig. 2 makes this argument clear. In Fig. 2 the incident beam is represented by the wave vector ko. This beam is initially diffracted by the lattice planes (hkl) which are represented by the reciprocal lattice vector r(hk1) The reciprocal lattice point (hkl) of course lies on the Ewald sphere in order that the Bragg condition be satisfied. The initially diffracted beam is represented by the wave vector kl. If now kl is rediffracted by a second set of planes (h'k'l') resulting in k2 the lattice point (h"k"l") must lie on the Ewald sphere if the double diffracted beam k2 is to be observed. It is clear that (h"k"l") is the vector sum of (hkl) and (h'k'l'), and that (h'k'l') need not in fact lie on the Ewald sphere. In the usual1,2 analysis the origin (000) of the reciprocal lattice is shifted to the point (hkl). However the translation (000)— (hkl) automatically shifts the point (h'k'l') to the Ewald sphere, i.e., (h'k'l')— (h"k"1"). It is to be seen that (hkl), the source reflection, must lie on the Ewald sphere; if it does not the translation (000) — (hkl) places the "new" origin off the sphere, not satisfying the Ewald construction and hence the Bragg condition. The point (h"k"1") of course must lie on the sphere if it corresponds to an observed doubly diffracted beam. It is to be noted that crystals having the hcp structure, as just discussed, are intrinsically capable (i.e., without lattice defects such as twins) of double diffraction, as contrasted to fcc and bcc crystals which are extrinsically capable (requiring lattice defects or reoriented lattice regions). The "allowed" bcc reciprocal lattice points (which comprise an fcc lattice) are (110), (2001, (211), (220), (310), (2221, (321), (4001, and so forth. The combination

Jan 1, 1967

By G. H. Gronhovd, W. H. Oppelt

Complete gasification to produce synthesis gas for manufacture of synthetic chemicals and fuels offers possibilities for utilization of the vast lignite fuel deposits in the U.S. The design of a slugging, fixed-bed, pressure gasification pilot Plant is described. Results are presented for initial operation of the plant at moderate pressure and using low-temperature char as fuel. Gasification of solid fuels is becoming increasingly important in the future overall energy picture, both in this country and abroad. Snythesis gas from gasification of coal consists chiefly of carbon monoxide and hydrogen, and after suitable purification and modification this gas can be used in the production of synthetic ammonia, methanol, higher alcohols, and synthetic liquid fuels, or as a reducing gas in the metallurgical field.' Most synthesis gas used in this country is made from natural gas, petroleum oils, and refining tail-gases, but the direct gasification of coal will become more attractive as costs of petroleum oils and natural gas increase.' A gasification process designed for use with brown coal and lignite was developed by the German Lurgi Co. prior to World War 11. In this process a fixed bed of lump-sized fuel is gasified with steam and oxygen at pressures ranging from 370 to 440 psi. Effective heat exchange between the ascending high-temperature gases and the descending solid fuel results in low gas outlet temperatures, high thermal efficiency, and favorable oxygen consumption. Lurgi gasifiers are used commercially for the production of synthesis gas or as a substitute for coal gas in many countries, including Germany, Czechslovakia, Great Britain, South Africa, Australia and Pakistan.3-6 The chief disadvantage of the conventional Lurgi gasifier which operates under nonclinkering conditions is that large quantities of excess steam must be supplied with the oxygen-steam mixture to reduce the temperature in the fuel bed below fusion temperature of the ash. Most of the steam leaves the gasifier undecomposed with the product gas and, after cooling, produces large amounts of dilute liquor. Evaluation of fixed-bed gasification experiments using dry-ash removal indicated that by increasing the oxygen concentration in the oxygen-steam blast the temperature in the combustion zone could be raised high enough to operate under slagging conditions. 7-8 Operation of a fixed-bed pressure gasifier under slagging conditions would reduce the steam requirements and the volume of waste liquor and would increase the steam decomposition, the overall thermal efficiency of the process, and the rate of gas production per unit cross sectional area of the gasifier. Because the Lurgi gasification process, modified for slagging operation, offered possibilities for efficient gasification of American lignite, the Bureau of Mines decided to investigate the process on a pilot-plant scale at the Grand Forks, N. D., Lignite Research Laboratory. DESIGN Description of Pilot Plant: A flowsheet of the Grand Forks slagging gasification pilot plant is presented in Fig. 1. The gasifier unit consits of the coal lock, the gas generator, and the slag lock. The fuel, which is periodically charged to the coal lock, moves by gravity flow into the generator and is continuously gasified by an oxygen-steam mixture introduced at the bottom of the gasifier. Molten slag is formed at the hearth and flows through a central taphole into a water quench bath. The slag-water slurry is periodically discharged from the slag lock. Process steam is generated and superheated in separately-fired, oil-burning units. Oxygen is supplied to the gasifier from a gas storage bank located some distance from the pilot plant. The capacity of the oxygen storage bank is about 130,000 standard cu ft at 2200 psig. Liquid oxygen is delivered by truck and is converted to gas during the filling operation. The superheated steam and oxygen are metered separately, then mixed and fed to the gasifier tuyeres. The gas generator is surrounded by a water jacket. The steam formed is separated in the steam drum and then discharged to the gas offtake, thus equaliz-

Jan 1, 1961

By R. B. Booth, W. F. Linke

The continuous interest of the American Cyanamid Company in producing superior polymeric flocculants and dispersants for the mining industry has resulted in a broad, general study of the physical chemistry of mineral-polymer systems. The aim of this program has been to establish a sound theoretical basis on which practical problems could be analyzed, and which would lead to the development of new and improved flocculating and dispersing agents. This study has required the cooperative effort of physical, analytical, and polymer chemists, as well as engineers familiar with mining problems. This report summarizes the results of some of our investigations over the past two years. Further studies along these and other lines are in progress and will be reported in the near future. MEASURING FLOCCULATION IN THE LABORATORY The foremost problem in studying flocculation in the laboratory has been to establish reproducible methods of comparing and evaluating the effectiveness of various reagents. Previous workers have measured settling -rates 1,8 filtration rates2 , and sediment volumes3 . Direct observations of the floc size, or the clarity of the supernatant liquid, can also be used. AIthough these parameters are all interrelated, they do not always measure the same physical properties of the suspensions, and hence are not equally useful or reliable. Whatever the method of comparison selected. the testing procedure must be carefully defined and standardized. The results must be satisfactorily reproducible, and for practical purposes must accurately reflect the results obtained in mill use. In order to compare the data obtained in these various types of test, the physical nature of the flocculated system must be examined in some detail. In an unflocculated system, only the larger or denser particles (the "sands") settle at a useful rate. The sediment can be observed to build up from the bottom of the cylinder, while the fines remain in suspension for a long period of time. If the system is not disturbed, even the microscopic fines can be seen to classify themselves as they fall over a period of days. When a high molecular weight polymer is added to the system, the colloidal particles are tied together randomly in flocs, and the sands and fines settle together. The line of demarkation between solid and liquid phases moves downward from the top of the cylinder. The settling rate of the flocs depends on their size and on their effective density. For a given mineral sample, the larger the flocs, the faster the settling. Since the large flocs cannot pack as closely as small ones, they produce a sediment with more pore spaces, and hence of larger volume, which is easier to pump and to agitate. Although it is nearly impossible to measure individual floc sizes, it is relatively easy to determine settling rates and sediment volumes. Of these two, the settling rate is simpIer to measure, and hence generally more practical. Reproducible sediment volumes can only be obtained by allowing the solids to pack for a rather long period of time (e.g., hours to days), and although the volumes thus observed are directly related to the degree of flocculation (average floc size), it is often impractical to wait so long for the results.

Jan 1, 1961

By Coopers &. Lybrand

INTRODUCTION From a tax viewpoint, United States mining companies are at a substantial disadvantage in the international mining area compared with companies in other major capital-exporting countries, according to a 1975 study by Coopers & Lybrand for the American Mining Congress. The study compared the relative tax position of a United States investor with investors from Belgium, Canada, France, Germany, Japan, the Netherlands, Switzerland, and the United Kingdom. The comparison was based on average rates of return realizable by investor companies on four mine-model investments in twelve capital-importing countries. The countries in which these investments were assumed to be made are varied in their tax systems so that a wide range of possibilities was included in the analyses. Each mine model was standard so that the rates of return for each investment by each investor were affected solely by the respective tax systems of the capital-exporting and capital-importing countries. UNITED STATES POSITION The study found that the United States competitive position, in terms of the average rate of return on an overall basis, ranked next to last among the nine capital-exporting countries in the study. Further, the United States overall rate of return is more than 20% below that of the country (Japan) with the highest overall rate of return. On an average return-on-equity basis, the United States' position among the nine investor countries, for each of the mine investments studied, was: copper, eighth; iron ore, sixth; nickel, fifth; manganese, a tie for last place. Figure 1 shows the overall results realized by each capital-exporting country on the basis of average return on equity. [ ] The frequency with which a given capital-exporting country attains the highest or next highest rate of return in a capital-importing country provides further insight and another basis for comparing the tax laws of the capital-exporting countries. Of the situations compared in the study, the United States never attained the highest rate of return, and reached the second and third highest rate of return in only two instances. The tables showing the frequencies of rankings are given in Table 1 for each of the four mine models. In summary, the United States mining company does not have a competitive advantage under the tax rules when competing for mineral concessions in other countries. The United States company, in fact, appears to be at a clear disadvantage. The study also considered the effect of the following recent tax proposals on United States investors in international mining ventures:

Jan 1, 1985

By L. Casagrande, C. O. Brawner, D. L. Pentz, B. Kennedy, E. Hoek, H. Q. Golder, K. Barron

H. L. HAMMERSLEY, B. C. Dept. of Mines. I would like to ask Mr. Brawner if the direction that he measured in the boreholes with the camera was referenced to magnetic north? MR. BRAWNER. Yes, it was. However, there is one problem that you may encounter if the ore body is magnetic; the borehole compass will give erroneous results. Three procedures are available to overcome this. The first one involves the use of a gyro compass instead of a magnetic compass. The second technique developed by Christiansen in Salt Lake City, is to orient the core with the use of scribing knives. The third method is to remove the compass needle and place a steel ball in the compass case. The hole must be drilled at a slight angle off the vertical. The location of the steel ball, which will roll to the underside of the hole is photographed. A borehole directional survey is then made, from which the bearing of the steel ball relative to the axis of the compass can be determined. The joint orientation is then related to this bearing at all depths. J. ROBINSON, B. C. Dept. of Mines. It was stated this morning that there have been some recent developments that give us the ability to drill horizontal drainage holes and install drainage pipes more cheaply than in the past. Would Mr. Brawner tell us some of the new developments and methods of installing these horizontal drain holes? MR. BRAWNER There are two zones in which you may be concerned about installing drains; the first is in the overburden soil, the second is in the rock. Each area requires a different installation method. If the rock is fairly massive and the holes will remain open, the simplest procedure is to use standard air track equipment or a percussion drill. At one of the mines in the Highland Valley, drains up to 300 ft. long have been installed with a percussion drill at a rate of one drain per day. The costs were in the order of $2.30 and $2.40 per foot. If the rock is quite soft, such as in a serpentine, it will be necessary to install a casing in the hole. The least expensive casing that will not deteriorate within the life required should be used. Plastic pipe or black iron pipe are commonly used. In soft rock, it may be necessary to drill with a casing to prevent the col-

Jan 1, 1971

By William Clark

A bean placed at the bottom of tubing in flowing wells is not a new idea. In fact, a device which in effect was a bottom bean was patented prior to 1890 by John D. Rockefeller. Because of the limited use of the bottom bean, prior to 1922, when it was installed in a few wells in the Santa Fe Springs field, California, the knowledge of the results of its operation is very meager. Since then various oil operators have experimented with bottom beans, but most of the records of the results obtained by their use are unavailable. For the past year or two, more attention has been directed to this method of producing wells; the theory and best means of application have been studied, and accurate records maintained. In theory, the use of bottom beans in flowing or gas-lift wells should secure certain advantageous results. With the exception of flow friction of the tubing and flow lines, there is no back-pressure held on the rising column of fluid. Thus severe beaning is possible with no danger of killing the well by too heavy a fluid column, a condition that often exists when a surface bean is used. The whole energy of the gas expanding to atmospheric pressure is utilized to lift the fluid in the well. The same effect is obtained when flowing a well through open tubing or casing. In using a surface bean this energy of expanding gas is wasted into the flow lines beyond the bean, because the fluid has reached the surface by the time the restriction is reached and the work of the gas performed. The bottom bean affords an efficient use of the expanding force of gas, utilizing less gas to raise the same amount of oil, hence allowing for lower gas-oil ratios. The flow restriction being at the entrance to the flow string, a more uniform mixture of gas and oil enters the tubing and a steadier backpressure is held on the well. The uniform mixture and high velocity attained while flowing through the bean, which is above the critical velocity required to suspend a spray condition of gas and oil, prevent slippage of gas, surging and heading with its resultant strain on the casing. The advantageous conditions allow positive control of production in gas-lift wells within limits set by the bean size and amount of input gas; and probably help to prevent sanding up of a well and the coning

Jan 1, 1931

By William M. Shakespeare

Thin sheet copper in wide widths and long lengths is being made electrolytically on a commercial scale at the Raritan Copper Works, Perth Amboy, N. J. This achievement is the result of three years of intensive development preceded by almost ten years of intermittent research. While the idea of producing sheet copper by the electrolytic process is old, the literature on the subject is limited. It is evident, however, that the idea has been attractive to many investigators for years, if the issued patents in the field may be taken as an indication. These patents may be divided into three groups. In one group metal is deposited on some form of endless cathode, of which a revolving cylinder or drum is the most common. The cathode moves slowly and at such a rate that metal of the desired thickness is deposited during one pass through the electrolyte. The deposited metal is stripped off the cathode at a point above the solution level and, after washing and drying, is wound up immediately. In the next group of patents a band of thin metal is increased in thickness by being guided through an electrolytic tank in opposition to suitable anodes. The third group covers means of producing sheets of definite dimensions where flat plates or revolving cylinders are used as cathodes. With such cathodes, sheets of definite dimensions are built up to the desired thickness and then stripped off. The process as worked out at the Raritan Copper Works combines the methods of the first two groups. In the first stage a sheet weighing 1 oz. to the square foot is produced on a revolving cylinder. This sheet is increased to any desired thickness in the second stage. In the first stage (Fig. 1) a lead-covered copper drum is mounted horizontally on a lead-lined concrete tank so that it. is partly immersed in the electrolyte. The cylindrical surface of the drum immersed in the electrolyte is opposed by lead anodes spaced about 36 in. from it. The electrolyte, made up of appropriate quantities of copper and sulfuric acid, flows into and out of the tank continuously. The ebullition of gas at the surface of the anodes induces a vigorous circulation of electrolyte between the anodes and the drum surface, which is augmented by means of compressed air introduced through a pipe between the ends of the anodes under the center of the drum, not indicated on the diagram. This

Jan 1, 1934

By R. F. Janke, J. F. Bossler

Introduction The commercial use of ion exchange resins to recover uranium evolved in the decade following 1950 when significant efforts were made to recover this vital element economically and efficiently from low grade ores. It was discovered that in both acid and alkaline systems, uranium exists as an anionic complex. This led to the application of anion exchange resins to concentrate and purify the uranium product. While this is the simplest statement of what an ion exchange resin does in the recovery of uranium, it should not be forgotten that the industry is continuing to evolve new methods in mining and processing which affect ion exchange operation. The inception of solution mining, the use of expanded bed ion exchange columns and the development of continuous ion exchange systems represent several examples of this continuing evolution in the industry. Throughout all of the changes in equipment designs and variations in solution compositions, the purpose of the ion exchange resin remains the same - to concentrate and purify uranium. Ion Exchange Resins When evaluating ion exchange resins for a uranium recovery project, the task that faces the chemist or process metallurgist is an assessment of the efficiency and, therefore, the economics of the various candidate resins for a given project. This assessment should include consideration of both physical factors such as particle size, density, hydraulic expansion and pressure drop as well as chemical factors such as the uranium loading and elution characteristics. For any given application, the differences in the physical and chemical behavior of ion exchange resins are a function of the conditions used to manufacture those resins. Therefore, an understanding of the way in which manufacturing variables affect resin behavior is important. Since most uranium recovery projects are concerned with recovery of the anionic complexes of uranium using anion exchange resins, this discussion will be restricted solely to the family of anion resins. The majority of the ion exchange resins used in commercial recovery operations look similar to the resins shown in Figure #1. [ ] Most of these materials are based on a styrene- divinylbenzene copolymer. The liquid monomers, styrene and divinylbenzene, are charged to a co-polymer reactor containing water. The water insoluble monomers are then mixed with the water to suspend oil-like droplets of the monomer mixture in. the water. The entire reaction mixture is then heated causing the monomers to polymerize into sol id thermoset plastic spheres. The spherical nature of the particle, the particle size distribution and the primary crosslinking level are fixed in this step in the production process. Figure 2 is a simplified illustration of this step of the process. [ ]

Jan 1, 1979