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By J. C. Rengel

Why and how the nuclear industry entered the electric power generation business is discussed in terms that nuclear energy was an undoubtedly additional energy resource and that it had promise of becoming a virtually inexhaustible generating power if breeder reactors could be devised. The growth of nuclear power, its total generation capacity, are related to the need of increasing energy requirements. The use of atomic energy for power generation will grow if it can compete economically with other fuels. The rate of growth, the time scale and direction of technical development is dependent on ore supplies. It is predicted that by 1980, 50% of all new power installations will be nuclear stations. INTRODUCTION Ten years ago, the nuclear industry was virtually unborn, the Atomic Energy Act of 1954 had just been signed, and peaceful applications of atomic power was little more than a phrase. Serious discussion of the practical application of nuclear energy to electric utility installations in the West was highly unlikely. The questions: Why and how did the nuclear industry enter the electric power generation business at all?; What is the predicted growth of nuclear power?; What technological changes do we predict?; How will the above changes affect the utilities? will in part be answered by this paper. WHY NUCLEAR POWER? The only justification for the West as for any other section of the country for atomic power is economics. Whatever energy source will provide the lowest generation cost will prevail; the only electric utilities which have shown an interest in nuclear power are those to whom it means lower generation costs. It is true that the Government sponsored the development of this technology long before there was any evidence of better economics; but a Government must take a very long-range view and it is the only type of organization that can afford to look more than twenty or thirty years into the future. The Government looked at projections such as shown in Fig. 1, which shows various rates of energy consumption, and matched these against estimates of fossil fuel resources. This showed that we will exhaust our low-cost fossil fuel supplies within 75 to 100 years and our total known fuel supplies within 150 to 200 years. The decision to invest in nuclear energy rested, therefore, on two arguments: a) It was undoubtedly an additional energy resource; and b) It has promise of becoming a virtually inex-haustable resource if breeder reactors could be devised if, while generating power, the reactors would produce more fissile materials than they consumed. This decision produced an unexpected bonus: nuclear power became competitive with fossil power much earlier than had been expected and, as shown in Fig. 2, forced the coal industry to take a hard look at its own position. I believe that the drop of coal prices shown here is the direct result of the entry of nuclear power into the energy supply business. At this point, it ought to be explained why and under what conditions nuclear power can be cheaper than fossil power. Fossil fuel cost to the utility is composed of two factors: a) Mining; and b) Transportation. As a result, the cost of fossil fuel to the utilities varies from 126, 106 Btu to 306, 106 Btu and higher. Nuclear energy is highly concentrated and the transportation factor becomes insignificant. Today the cost of nuclear fuel to the utility ranges between 156

Jan 1, 1967

By G. R. Fitterer

Trends in slag composition in acid open hearth practice, particularly the variation in iron and manganese oxides during refining, are reviewed. A procedure which is currently being used to control the refining reactions is described and partially explained through slag phase diagram considerations. In addition, the ionic nature of these slags is considered and it is proposed that silica combines with the oxide ion and forms a series of complex silicate ions. The effect of basic oxides such as FeO, MnO, and CaO is apparently additive in furnishing the oxide ion for this purpose. IT is now possible to control192 the refining of steel in an acid open hearth furnace to a surprising degree. This control has been brought about through the knowledge and use of several facts and tests. These are: 1—A planned charge analysis based on a knowledge of the type of fuel, the bath depth, etc. 2—Knowledge and control of rate of fuel input as well as its method of atomization, if any. 3—Determination of the slag fluidity. 4—Bath temperatures (use of the Pt-Pt-Rh thermocouple). 5—Periodic determination of carbon content (use of the Carb-analyzer or Carbometer). Through the use of these control methods, a heat may be made to "melt-in" with or without residual manganese and silicon. The rate of refining or carbon elimination may be controlled accurately, and the desired tapping (or "go-ahead") carbon may be approached at a rapid or slow rate according to preference and type of steel being made. Thus operating time may be predetermined and greatly reduced. Further, all of the refining can be done without any ore or other oxidizing additions, and the heat may be brought to the desired tapping temperature simultaneously with the attainment of the desired carbon analysis. This control procedure is being used in most acid plants in the United States today and has been de- scribed.', ' Although the slag metal relations are sufficiently well known to control the heat as stated, yet it is possible to account for only about two-thirds' of the 0, which must have been used for the elimination of the carbon and other metalloids. Phase diagram studies' of the slags, slag weights, the metal particle theory,' ore additions, limestone additions, etc., may all be taken into consideration; yet, the mechanism of oxidation and source of all the oxygen remain to be explained. This paper reviews some of the principles upon which the control methods were based and proposes other possibilities regarding the unexplained phenomena. Old and New Acid Open Hearth Practice Some of the recent statements in the literature regarding acid open hearth operation are based on old heat data which are not typical of modern American practice. These old heat logs have been responsible for many misconceptions regarding the acid hearth practice and represent heats which do not coincide with modern methods. In recent years, American practice has been altered materially and these changes will be briefly described. Modern American Practice: A typical log of a modern American acid open hearth heat' is shown in Fig. 1 which indicates that the refining time was only two hours, during which the carbon was eliminated rapidly and the bath temperature increased at a fast rate. The manganese and silicon contents of the metal were essentially eliminated and their oxides had become constituent parts of the slag at the "melt-down" period. During the course of the heat, the MnO content of the slag remained practi-

Jan 1, 1954

By Sidney A. Swensrud

The 24-in. crude-oil line. and- the 20-in. petroleum-products line, built as a war emergency by the United States Government and running from points in Texas to the New York-New Jersey-Philadelphia area, were constructed only because of the inability during the war to continue nor-inal ocean tanker movements from the Gulf Coast to the East Coast. The postwar disposition of these lines may be considered as one of the important problems of the postwar readjustments in the oil industry, if not the most important one. In the first place there appear to be many complications in keeping the lines in crude-oil or petroleum-products service. This conclusion was reached also in a study by T. E. Swigart, President of Shell Pipeline Corporation.' Since the lines are too largc for any one company to utilize, and since it appears likely that a surplus rather than a deficiency of tankers will exist after the war, most companies probably will feel that they can achieve lower costs and more satisfactoiy results from the resumption of tanker movements to the East Coast rather than by endeavoring to take over or utilize the capacity in the large pipe lines under some tariff-rate structure. As for using the lines for making deliveries to the middle west or central eastern areas, it would appear that when the war is over and normal transportation arrangements can be resumed, there will be no need of any extra capacity for such deliveries. Therefore, the War Emergency lines could be utilized for this purpose only if privately owned pipe lines and other transportation facilities were to be displaced. Against this rather unprepossessing prospect for use in petroleum or product service is the fact that at or near the source of the lines lie the greatest developed gas reserves in the world, for much of which there is at present no satisfactory market outlet; and at the other end of the lines lies one of the greatest concentrations of population in the world, which is not now served by natural gas. These who have been fortunate enough to use natural gas can have little doubt of its desirability as a domestic fuel. An analysis of the market potentialities and comparison with other areas where natural gas is being distributed, and where it has displaced manufactured gas, makes it seem clear beyond question that the full capacity of both of the large-diameter lines could be fully utilized in this service, and this without creating any undue drain upon the gas reserves of the nation. On the basis of an estimated capacity of roo to 120 billion cubic feet per year, the gas that might be marketed through these lines would amount to only 3 to 4 per cent of the total con-

Jan 1, 1944

By Sidney A. Swensrud

The 24-in. crude-oil line. and- the 20-in. petroleum-products line, built as a war emergency by the United States Government and running from points in Texas to the New York-New Jersey-Philadelphia area, were constructed only because of the inability during the war to continue nor-inal ocean tanker movements from the Gulf Coast to the East Coast. The postwar disposition of these lines may be considered as one of the important problems of the postwar readjustments in the oil industry, if not the most important one. In the first place there appear to be many complications in keeping the lines in crude-oil or petroleum-products service. This conclusion was reached also in a study by T. E. Swigart, President of Shell Pipeline Corporation.' Since the lines are too largc for any one company to utilize, and since it appears likely that a surplus rather than a deficiency of tankers will exist after the war, most companies probably will feel that they can achieve lower costs and more satisfactoiy results from the resumption of tanker movements to the East Coast rather than by endeavoring to take over or utilize the capacity in the large pipe lines under some tariff-rate structure. As for using the lines for making deliveries to the middle west or central eastern areas, it would appear that when the war is over and normal transportation arrangements can be resumed, there will be no need of any extra capacity for such deliveries. Therefore, the War Emergency lines could be utilized for this purpose only if privately owned pipe lines and other transportation facilities were to be displaced. Against this rather unprepossessing prospect for use in petroleum or product service is the fact that at or near the source of the lines lie the greatest developed gas reserves in the world, for much of which there is at present no satisfactory market outlet; and at the other end of the lines lies one of the greatest concentrations of population in the world, which is not now served by natural gas. These who have been fortunate enough to use natural gas can have little doubt of its desirability as a domestic fuel. An analysis of the market potentialities and comparison with other areas where natural gas is being distributed, and where it has displaced manufactured gas, makes it seem clear beyond question that the full capacity of both of the large-diameter lines could be fully utilized in this service, and this without creating any undue drain upon the gas reserves of the nation. On the basis of an estimated capacity of roo to 120 billion cubic feet per year, the gas that might be marketed through these lines would amount to only 3 to 4 per cent of the total con-

Jan 1, 1944

By J. D. Forrester, Erich Sarapuu, T. C. Cheasley

Electrocarbonization of coal will produce cheap fuel for electric power plants. The complete electrification of industry and domestic power consumption of gas generated as fuel in situ is a worthwhile goal for the coal industry. Production of synthetic oils, gasoline, and coal chemicals from coal gas will create additional new markets for the coal industry. RANKING high in the nation's wealth of natural resources, readily available liquid and gaseous hydrocarbons are nevertheless falling steadily in supply, while industrial requirements for hydrocarbon compounds are greater today than during the peak period of World War 11. Theoretically, several processes will increase the future supply of liquid and gaseous fuels, but the practical application of these methods depends on two controlling factors: 1—that enough carbon-bearing material be available to warrant industrial exploitation, and 2—that the given process be physically, chemically, and economically applicable. With regard to the first factor, the Western Hemisphere is fortunately well supplied, see Table I. Table I not only shows the extensive occurrence of many fuel sources, but also the wide discrepancy among quantities of fuel substances in various materials. Great quantities of potentially available fuel have not been recovered. Many coal seams and other hydrocarbon-bearing deposits are so situated that production is uneconomical by present-day mining or oilwell pumping methods, and even in many exploitable deposits where the most efficient production methods are used a large proportion of material is never recovered. Industrial production of hydrocarbon liquids and gases must be better balanced, and processes must be developed which will convert coals, oil sand, and oil shales into fabricated commodities. Among the techniques employed by scientists and engineers to secure fuel masses not economically available at present are several processes of underground gasification. These are all concerned with converting raw materials into gaseous products by thermal treatment of source substances while they are still in the ground and therefore require that a subsurface heat (fire) zone be induced to affect pervasively the buried raw fuel deposit. This is possible only where sufficiently high temperature exists and where the fuel material is satisfactorily opened or channelized. Also, to compete successfully with common mining or oilfield practices, a given method must be controllable and must not involve greater operational expense than regular methods. Preferably, it should be less expensive. Since the close of World War 11, England has continued field studies of underground gasification methods. Experiments have also been conducted in Sweden, Estonia, France, Belgium, Italy, and Germany. The U. S. Government has sponsored field studies at Gorgas, Ala., and because gasification in situ is vitally important to establishment of the nation's fuel resources, the mining engineering department at the Missouri School of Mines has been actively interested in developing the process of underground electrocarbonization. Research was conceived by Erich Sarapuu, and preliminary experiments in treating bituminous coal were begun at the school in September 1947. The work consisted of laboratory experimentation until June 1948, at which time, through a cooperative agreement with the Sinclair Coal Co. of Missouri, pilot plant studies were instituted at the Tiger mine of the Hume-Sinclair Coal Co. near Hume, Mo., to test the encouraging laboratory findings. This comprehensive field program was pursued continuously until the beginning of 1952."'" The broad term underground electrocarbonization describes a unique gasification process that releases by electricity the desired economic constituents from a raw carbon-bearing material, converting the substance of such deposits into deliverable form and greatly reducing the cost-of-mining factor. Underground electrocarbonization is not subject to the major limitations of many other underground gasifi-

Jan 1, 1955

By William A. Mitchell

AN X-ray diffraction pattern for "calcined cold precipitated ferric oxide" is reproduced diagrammatically along with data for other iron oxides by R. C. Mackenzie.1 This pattern, which shows spacings higher than those for any other form of iron oxide, was obtained from material prepared and photographed at several different times over a period of weeks, but later attempts to reproduce the results were unsuccessful the pattern of hematite being obtained every time.2

Jan 9, 1953

By E. W. Davis

About a year ago several carloads of magnetite ore from New York state were sent to the Mines Experiment Station at the University of Minnesota for grinding and concentration tests. The flowsheet used minus ½-in. ore as feed to a 3 by 3-ft. overflow cylindrical ball mill operating at 37 r.p.m. in closed ciruit with a vibrating screen. The screen oversize was returned to the mill for further grinding and the screen undersize was concentrated magnetically. Several different screens, between 8 and 35 mesh, were used and it was desired to determine the capacity of the ball mill operating in closed circuit with each screen. From earlier tests using this flowsheet and equipment, it was known that when comparatively coarse screens were used it was necessary to increase the grinding dilution in the mill beyond that ordinarily used when grinding in closed circuit with classifiers. It was also known that after the circulating load had passed about 100 per cent of the original feed it tended to increase rapidly in amount, and grinding conditions appeared to become unstable. These conditions may be peculiar to this small ball mill but they were very marked in this series of tests. If conditions were not right, the mill would suddenly "go dead," and it was found desirable to operate so that the occasional rattle of balls against the lining of the mill could be detected. As the test proceeded, it was found that the cure for almost any difficulty that developed in the grinding circuit was to add more water to the mill feed. To begin with, the mill was operated at pulp densities of from 60 to 65 per cent solids but gradually it was found that with higher dilutions grinding conditions were more stable and the ball-mill power per ton of screen undersize decreased. Actually, the most satisfactory operating conditions were secured with a pulp density of about 40 per cent solids. The ore was hard but it was possible to grind ½-in. feed so that all would pass an 8-mesh screen with a power expenditure in the ball mill of less than 2½ kw-hr. per ton and with a circulating load of less than 50 per cent. With so high a dilution in the feed to the mill, it was feared that steel wear would be very high. However, the mill was not noisy, and in order to find exactly the grinding conditions, a special test was made at the end of the main series of tests. After the mill had been in operation for several hours and grinding conditions had become balanced, the power to all the equipment was shut off at a given signal. The ball mill made about one revolution after the power was turned off, and just as soon as the mill stopped, a spout was placed at the discharge trunnion and all of the pulp that flowed out of the mill was caught in tanks. The manhole in the mill was then opened and the whole mill charge dumped into other tanks. Both of these products were weighed, dried, and weighed again. The balls were screened out and weighed, and the dry solids were sized. In this way it was possible to determine the nature of the pulp and the ore within the mill under actual grinding conditions.

Jan 1, 1947

By E. W. Davis

About a year ago several carloads of magnetite ore from New York state were sent to the Mines Experiment Station at the University of Minnesota for grinding and concentration tests. The flowsheet used minus ½-in. ore as feed to a 3 by 3-ft. overflow cylindrical ball mill operating at 37 r.p.m. in closed ciruit with a vibrating screen. The screen oversize was returned to the mill for further grinding and the screen undersize was concentrated magnetically. Several different screens, between 8 and 35 mesh, were used and it was desired to determine the capacity of the ball mill operating in closed circuit with each screen. From earlier tests using this flowsheet and equipment, it was known that when comparatively coarse screens were used it was necessary to increase the grinding dilution in the mill beyond that ordinarily used when grinding in closed circuit with classifiers. It was also known that after the circulating load had passed about 100 per cent of the original feed it tended to increase rapidly in amount, and grinding conditions appeared to become unstable. These conditions may be peculiar to this small ball mill but they were very marked in this series of tests. If conditions were not right, the mill would suddenly "go dead," and it was found desirable to operate so that the occasional rattle of balls against the lining of the mill could be detected. As the test proceeded, it was found that the cure for almost any difficulty that developed in the grinding circuit was to add more water to the mill feed. To begin with, the mill was operated at pulp densities of from 60 to 65 per cent solids but gradually it was found that with higher dilutions grinding conditions were more stable and the ball-mill power per ton of screen undersize decreased. Actually, the most satisfactory operating conditions were secured with a pulp density of about 40 per cent solids. The ore was hard but it was possible to grind ½-in. feed so that all would pass an 8-mesh screen with a power expenditure in the ball mill of less than 2½ kw-hr. per ton and with a circulating load of less than 50 per cent. With so high a dilution in the feed to the mill, it was feared that steel wear would be very high. However, the mill was not noisy, and in order to find exactly the grinding conditions, a special test was made at the end of the main series of tests. After the mill had been in operation for several hours and grinding conditions had become balanced, the power to all the equipment was shut off at a given signal. The ball mill made about one revolution after the power was turned off, and just as soon as the mill stopped, a spout was placed at the discharge trunnion and all of the pulp that flowed out of the mill was caught in tanks. The manhole in the mill was then opened and the whole mill charge dumped into other tanks. Both of these products were weighed, dried, and weighed again. The balls were screened out and weighed, and the dry solids were sized. In this way it was possible to determine the nature of the pulp and the ore within the mill under actual grinding conditions.

Jan 1, 1947

By A. U. Seybolt

The solubility of oxygen in a iron has been determined in the range between 700° and 900°C. The solubility is a function of temperature and varies from about 0.008 pct oxygen at 700°C to atureandabout 0.03 pct at 900°C. The heat of solution is approximately +15,500 cal per mol. AS pointed out in a recent paper by Kitchener et al.,1 there has been a lack of agreement among many investigators even as to the order of magnitude of the solid solubility of oxygen in the various forms of iron. This lack of agreement is attributable in large part to the difficulties in the determination of a small oxygen solubility; but because the problem has remained so long unsettled, it also indicates a lack of interest which is rather surprising when the demonstrated importance of small amounts of soluble nonmetallic impurities in iron is considered. The work of Kitchener et al. apparently leaves the solubility of oxygen in iron in a satisfactory state, but no attempt was made to investigate the solubility in a iron. The solubility of oxygen in a iron is actually of greater interest, since it is in this form that iron and mild steels are employed ordinarily. That the effect of oxygen in iron is of more than theoretical interest has been well established by Fast,' and more recently by Rees and Hopkins," who demonstrated that oxygen in the range between 0.0008 and 0.27 wt pct has a pronounced effect upon the mechanical properties. Previous Work and Methods Used To report in detail the literature relating to the solubility of oxygen in a iron would require an inordinate amount of space. For those interested in reviewing this work, a bibliography4-13 of the more significant papers is appended. In general, two methods of studying this problem have been used. One is the gas-metal equilibrium method where the H2O-H2-Fe or the CO2-CO-Fe equilibria have been used. The other is the more direct approach of the oxidation of thin strips of pure iron by packing in mill scale or by air or gaseous oxygen at some desired temperature. In this method oxygen is allowed to oxidize the surface and then to diffuse inward until saturation is obtained. In the gas-metal equilibrium method the oxygen dissolved in solid iron at a given temperature is proportional to the ratio of the water vapor-hydrogen pressures or the CO2-CO pressures over the sample for small ratio values. If the ratio becomes higher than a critical value, then an oxide phase makes its appearance (the solution becomes supersaturated). In principle, it is possible to use a series of H2O/H2 or CO2/CO ratios and to find by analysis the corresponding amounts of oxygen in solid solution at constant temperature. At the point where a very small increase in the gas ratio (increase in oxidizing powder) produces a large increase in oxygen content, the solid solubility limit is reached. Alternately, if the critical ratio is known, it is possible to use the procedure of Kitchener et al.1 and to use a ratio which is near but below the critical one. The solubility corresponding to this lower ratio will not be the saturation solubility at the temperature employed, but the saturation solubility can be calculated by multiplying the measured solubility by the critical ratio over the ratio used. However, in the case where oxygen gas is the oxidizing medium, the saturation solubility is not a function of pressure, providing the pressure exceeds the dissociation pressure of FeO in equilibrium with iron. This is 1.2x10-16 mm at 800 °C, according to Dushman." As pointed out by Darken17 in discussing the FeO phase diagram, most of the possible errors tend to yield high values of oxygen solubility. For example, one circumstance which evidently caused the reporting of many high values was the use of finely divided or powdered iron in the gas equilibrium method. Because of the large surface area of such a sample, and the likelihood of some surface contamination if only by exposure to air, the results tended to be high. The direct oxidation method which was used in this work has the advantage in that it is simple and direct, but it suffers from one disadvantage: equilibrium can only be approached from the low oxygen side. The important factors to be kept under control are the following: 1—use of high purity iron to avoid internal oxidation (oxidation of readily

Jan 1, 1955

By Joseph D. McClung, L. G. Austin

INTRODUCTION Run-of-mine (ROM) coal generally falls into two major groups, that from underground mining (continuous mining machines) and that from strip mining. Continuous miners produce a finer product and strip mining a coarser product that is crushed to produce the desired size. Further size reduction depends on the proposed use or uses for the coal. For example, for pulverized coal power plants as a rough general rule, the coal is ground or pulverized to 80% less than 200 mesh (0.075 mrn) before it is blown into the boiler furnace as a coal-air suspension. For solvent-refined coal processes, the desired size is about <30 mesh ( <0.60 mm) ; in fluidized bed uses, the coal is typically < in. ( < 1.6 mm); for Lurgi pressure gasifiers, the desired size is 1-2 1/2 in. (25-63 mm). Coal preparation plants typically work with a coarse size fraction of 6x1 1/4 in. (15Ox32mm). It seems probable that the next decades will see much wider application of coal cleaning before use and that the stages of crushing and pulverization from ROM coal to end use will be integrated with coal cleaning processes. Since different cleaning processes work better with different ranges of coal size and the degree of cleaning possible is inti-

Jan 1, 1979

By Waldemar Gooskov

IN his well-known work,1 published in 1867, P. Ritter von Rittinger gives the following formula for the velocity of a particle settling in still water: v=C/D(a -1) and for average grains he assigns to the constant C a value of 2.44. As is well known, the value of C depends on the shape of the grains resulting from crushing, and it is obvious that values obtained for coal particles will not only be very different from those for ore particles but will even be characteristic of coal from a certain seam. The only way of finding an accurate value of C, applicable to conditions in Russia, where all coal-washing plants are built on German principles, is to conduct a series of tests of which the results will determine what value of C is to be used. In the investigation outlined in this paper the theories of Prof. Robert H. Richards2 were taken as the basis of the tests; the slimes were removed from the charges to eliminate the trouble of handling them; and to obtain accurate results free from errors inevitably resulting from human observation, photographic apparatus was used and a series of pictures taken which form the basis for constructing a suitable diagram of velocity vs. time.3

Jan 1, 1928

By K. A. Jackson, J. D. Hunt

A general theory for the growth of lamellar and rod eutectics is presented. These modes of growth depend on the interplay between the diffusion required for phase separation and the formation of interphase boundaries. The present analysis of these factors provides a justification for earlier approximate theovies. The conditions for stability of rod and Lanlellar structures are consitleved in terms of the mechanisms by which the structure can change. The mechanisms considered include both small adjustments to the lnnzellar spacing due to the motion of lamellar faults, and catastrophic changes due to instabilities. It is concluded that stable growth occurs at or near the minimum interface undevcooling for a gizierz growth rate. The conseqrlences of the existence of a diffusion boundary layer at the interface are discussed. The experimental results for the variation of growth rate, undercooling, and Lanzellar spacing are cornpared with the theory. We believe that the theory presented in this paper provides an adequate basis for understanding the complex phenomena of lanzellar and rod eutectic growth. The growth of lamellar eutectics has been the subject of several theoretical and many experimental studies. The foundations for the theoretical work were laid by zenerl and Brandt2 in their analyses of the growth of pearlite. Zener estimated the effect cf diffusion, and took into account the surface energy of the lamellar structure. He found that the lamellar structure could grow in a range of growth rates at a given undercooling provided the lamellar spacing was appropriate for the growth rate. Since pearlite grows with only one growth rate and one lamellar spacing at a given undercooling, there is clearly an ambiguity in the theory. Zener removed this ambiguity by postulating that the growth rate was the maximum possible at the given undercooling. He predicted then that the product of the growth velocity v and the square of the lamellar spacing A should be constant, i.e., A2v = const. Brandt2 started out by assuming that the interface between the lamellae and austenite was sinusoidal. Because of this, the ambiguity encountered by Zener did not arise. Brandt was able to obtain an approximate solution to the diffusion equation, but, since he did not take into account the surface energy, his considerations are incomplete. Tiller3 applied some of these ideas to the growth of eutectics, and proposed a minimum undercooling condition to replace the maximum velocity condition used by Zener. These conditions are formally identical. Hillert4 extended the work of Zener. He found a solution to the diffusion equation assuming the interface to be plane. Taking surface energy into account, and applying Zener&apos;s maximum condition, he was able to calculate an approximate shape of the interface. Jackson et al.5 used an iterative method employing an electric analog to the diffusion problem to refine the calculation of interface shape. It was found that the interface shape calculated from a plane-interface solution to the diffusion equation was negligibly different from the exact solution. The method provided an analog only for eutectics for which the volumes of the two phases are equal, growing from a melt of exactly eu-tectic composition. There has also been considerable experimental work on eutectics, Several experimenters8-10 found that A2v is constant as predicted by Zener.1 Hunt and chilton10 demonstrated that ?T/v1/2 is also a constant for the Pb-Sn system as predicted. Lemkey et al.11have recently found A2v to be constant for a rod eutectic. In the present paper, we present the steady-state solution for the diffusion equation for a lamellar eutectic growing with a plane interface, for the general case, that is, with no restriction on the relative volumes of the two phases, and with the melt on or off eutectic composition. A similar solution is also found for a rod-type eutectic. Expressions are obtained for the average composition at the interface and the average curvature of the interface. These equations for the average composition and curvature are similar in form to those derived by Zener1 and Tiller,3 and provide a justification for some of the approximations made by these authors. The mechanisms by which the spacing in a lamellar structure can change are considered. The important mechanism for small changes in lamellar spacing involves a lamellar fault. Examination of the stability of lamellar faults leads to the conclusion that the growth occurs at or near the extremum.* The insta- bilities which can develop in a rodlike structure are also discussed, resulting in the conclusion that this structure also grows at or near the extremum. Comparison of the conditions for rod and lamellar growth permits a prediction of the surface-energy anisotropy required to produce rods or lamellae for various volume-fraction ratios. The diffusion equation predicts the existence of a diffusion boundary layer at the eutectic interface unless the eutectic has 0.5 volume fraction of each phase and is growing into a liquid of eutectic composition. This boundary layer is such as to make the composition in the liquid at the interface approximately equal to the eutectic composition. This boundary layer permits changes in composition during the zone refining of eutectics. Photographs of the eutectic interface of a growing transparent organic eutectic system have been made. Both the components of this eutectic are transparent organic compounds that freeze as metals do.12 The in-

Jan 1, 1967

By C. D. Rugen

Ground raw material as fed to the cement kiln generally is a mixture of two to four components, each of which may have widely varying physical and grinda-bility characteristics. Chemically similar materials found in the same or different deposits may also differ appreciably in physical properties, depending on their crystalline structure. The calcareous or principal cement raw material may vary from a soft chalky limestone or friable cement rock at one plant to a hard limestone or oyster shell at another plant. Argillaceous materials include clays, shales and granulated slag. Siliceous and/or iron-bearing material, including sand, iron ore and cinders, may also be required in addition for certain types of mixes. Frequently differences in grindability characteristics are not given the attention they deserve in comparing grinding results at one plant with those at another or in selecting suitable circuits or equipment. Many types of crushing and grinding units are used, with numerous open and closed circuit arrangements. Both wet and dry processes are used successfully. When rebuilding or installing a new grinding department, the choice of process or circuit in some cases may be dictated by the nature of the materials being ground but in other cases it may be the result of good salesmanship or a compromise. An over-all evaluation of wet process vs. dry process grinding should not be made without considering the effects on all other producing departments in the plant. In general, the more recent grinding installations, both wet and dry, have employed closed circuits, but many older open circuits are still in successful operation. In our experience, closed-circuit raw grinding pays, because of lower power consumption and less operating labor per ton of product. However, some dry open-circuit mills with improved equipment and operating methods compare favorably in performance with dry closed-circuit units. Character of Ground Raw Mix (Kiln Feed) Grinding in the cement plant in general requires a considerably her product than in most ore-grinding operations. A distinction between fineness of kiln feed and of cement may be made in that kiln feed preferably should be ground to a limiting size,whereas finished cement should include a range of size fractions including sufficient fine flours to give a product with satisfactory physical qualities. Grinding raw mix for kiln feed therefore is more comparable to ore grinding, although the limiting size is smaller. Most kiln feeds are ground from 85 to 95 per cent passing 200-mesh, depending on mix composition. Some cement chemists believe that raw materials cannot be too he, since finer particles theoretically combine more readily and intimately in the kiln. However, there are certain definite practical objections to extremely fine kiln feed, such as loss of kiln dust, unstable material flow through the kiln and greater dificulty in control of chemical composition due to dust loss, and the disproportion of fines from the more easily ground component. Our

Jan 1, 1947

By C. D. Rugen

Ground raw material as fed to the cement kiln generally is a mixture of two to four components, each of which may have widely varying physical and grinda-bility characteristics. Chemically similar materials found in the same or different deposits may also differ appreciably in physical properties, depending on their crystalline structure. The calcareous or principal cement raw material may vary from a soft chalky limestone or friable cement rock at one plant to a hard limestone or oyster shell at another plant. Argillaceous materials include clays, shales and granulated slag. Siliceous and/or iron-bearing material, including sand, iron ore and cinders, may also be required in addition for certain types of mixes. Frequently differences in grindability characteristics are not given the attention they deserve in comparing grinding results at one plant with those at another or in selecting suitable circuits or equipment. Many types of crushing and grinding units are used, with numerous open and closed circuit arrangements. Both wet and dry processes are used successfully. When rebuilding or installing a new grinding department, the choice of process or circuit in some cases may be dictated by the nature of the materials being ground but in other cases it may be the result of good salesmanship or a compromise. An over-all evaluation of wet process vs. dry process grinding should not be made without considering the effects on all other producing departments in the plant. In general, the more recent grinding installations, both wet and dry, have employed closed circuits, but many older open circuits are still in successful operation. In our experience, closed-circuit raw grinding pays, because of lower power consumption and less operating labor per ton of product. However, some dry open-circuit mills with improved equipment and operating methods compare favorably in performance with dry closed-circuit units. Character of Ground Raw Mix (Kiln Feed) Grinding in the cement plant in general requires a considerably her product than in most ore-grinding operations. A distinction between fineness of kiln feed and of cement may be made in that kiln feed preferably should be ground to a limiting size,whereas finished cement should include a range of size fractions including sufficient fine flours to give a product with satisfactory physical qualities. Grinding raw mix for kiln feed therefore is more comparable to ore grinding, although the limiting size is smaller. Most kiln feeds are ground from 85 to 95 per cent passing 200-mesh, depending on mix composition. Some cement chemists believe that raw materials cannot be too he, since finer particles theoretically combine more readily and intimately in the kiln. However, there are certain definite practical objections to extremely fine kiln feed, such as loss of kiln dust, unstable material flow through the kiln and greater dificulty in control of chemical composition due to dust loss, and the disproportion of fines from the more easily ground component. Our

Jan 1, 1947

By F. W. Jessen, G. F. Sears

Cavities in massive salt for the purpose of storage of liquid hydrocarbons have assumed a prominent position in recent years. This paper describes a program to facilitate leaching operations for the formation of specifically shaped storage cavities. Various forms and sizes of cavities may be possible through use of the techniques developed. INTRODUCTION The creation of large under ground storage facilities for natural gas and liquified petroleum gases has been practiced for many years. Use of cavities obtained through solution of salt for this purpose is a fairly new and novel approach, but has gained increasing importance due mainly to the economics of this type of mining operation opposed to hard rock mining or surface and pit storage.1 The idea of controlling the shape of any cavity dissolved from massive salt has not been a prime consideration of companies engaged in the formation of storage space. This has been due mainly to an insufficient knowledge of the mechanics of the leaching process and a dearth of published information dealing with both the desirability and ease of control possible for this type of operation. Two distinct advantages are readily ascertainable. From a stability standpoint (i.e., the ability of the completed cavity to withstand stress imposed by overburden pressure and lateral tectonic stresses) a controlled cavity may be generated which will yield the most favorable attitude to these external forces and remain in operable use for longer periods of time.2 A second advantage is that for a given volume, a sphere (which is one controlIed shape possible) represents the minimum surface area exposed. This may become more important in the future when dealing with refrigeration and product losses in underground cavities. The basic solution mining process, without regard for controlling the final shape, is quite simple in that the equipment and materials required to dissolve a cavity in a salt dome or layered salt section are a source of fresh water, a circulating pump, several strings of tubing and a means of disposal of the return brine. To add control measures involves the use of an inert blanket material above the position at which solution proceeds. Initially the annulus of the largest wash pipe is filled with this blanket down to the top of the proposed cavity. The bottom of the proposed cavity is determined by the depth of the original drilled hole or a blanking plug. The blanket material is added incrementally in stages and displaces the water in the enlarging cavity downward. Where the blanket material has displaced the water, no further solution takes place. The rate at which to add this controlling blanket material so as to be able to form specifically shaped cavities of any particular size is of considerable importance. THE PROBLEM It is desirable to have as much information as possible as to the behavior of the mining operation since visual observation obviously is impossible. It would be advantageous, for example, to have a step-by-step program showing how much salt is to be removed, at what rate and the time required. This type of program would facilitate procurement of surface equipment and provide for proper management of manpower requirements, as well as yield a host of intangible benefits. As mentioned earlier, only rule-of-thumb estimates were available until recently and made this type of pre-planned operation haphazardous at best. Recent work3-&apos; has done much to clear up the ambiguities and inaccuracies attendant to solution mining of salt and has helped to place the operation on a more scientific basis. This study attempts to correlate much of the information thus far developed into a complete mining program; to extend the idea of controlled solution mining to include not only spheres, but solid conic sections of the ellipsoidal variety; and to refine the computation of the rate of removal of salt during the various stages of mining. Dommers3 and IIemson4 made extensive laboratory

Jan 1, 1967

By Allen F. Agnew

Geologic studies for the Federal and State governments in the Wisconsin-Illinois-Iowa zinc-lead district were begun in 1835, and subse¬quent surveys were made in the three states at in¬tervals until 1916, see Table I. The current study of the district, Fig. 1, by the U. S. Geological Survey was begun in 1942 in the hope that a systematic investigation would help in¬crease production of zinc and lead, then in ex¬tremely short supply. Major emphasis of the USGS program was on detailed mapping of the geologic structure and ore deposits. Preliminary maps and reports covering localities of intensive mining ac¬tivity have been published during the course of this study, Fig. 2. By 1950 the objectives of the study by the USGS were revised; 7Y2-min quadrangles of relatively unprospected localities as well as inten¬sively mined localities are now being mapped in Wisconsin. Since 1945 the investigation in Wiscon¬sin has been made in cooperation with the Wiscon¬sin Geological and Natural History Survey, and since 1951 the geologic mapping of areas of particu¬lar interest in Iowa has been performed in coopera¬tion with the Iowa Geological Survey. At intervals since 1943 the Illinois State Geological Survey has mapped the geologic structure and ore deposits of localities in the Illinois part of the mining district, see Fig. 2.

Aug 1, 1955

By D. A. Vaughan

COMPOUNDS in the U-N system have received little attention since the work of Rundle, Baen-ziger, Wilson, and McDonald in 1948.&apos; They described the three nitrides UN2, U2,N2, and UN as the only phases in this system. The region between UN and U,N, was reported to consist of the two phases while the region between U,N, and UN, is a homogeneous solid solution in which the U,N3, structure transforms with increasing nitrogen content into the UN, structure. This report presents the results of an X-ray diffraction study of the phases present in the region of the diffusion gradient on uranium reacted with nitrogen at several temperatures between 500" and 915°C. Several samples of uranium used in U-N gas-metal reactions by Mallett and Gerds were available for X-ray diffraction study of phase composition through the diffusion gradient. X-ray diffraction data were obtained using filtered iron radiation at intervals through the diffusion layer of the several samples on filings or turnings at approximately 0.001 in. increments down to the base metal. Representative X-ray diffraction patterns from various layers through the case on uranium after nitriding at 850°C are shown in Fig. 1. These data show that three different phases are present in the

Jan 1, 1957

The following tables showing the output of coal during the year 1913 have been secured y R. H. Morris, of Pocahontas, Alberta, Canada, for the use of the Committee through the kindness of John T. Stirling, Chief Mine-Inspector of Alberta. The principal mining districts are, Crow&apos;s Nest Pass, Calgary, Lethbridge, and Edmonton, the outputs from which were as follows: Tons Crow&apos;s Nest Pass 1,849,435 Calgary...: 627,461 Lethbridge ; 966,020 Edmonton 863,430 Total 4,306,346 Of this quantity there was exported to the United States, principally from the Crow&apos;s Nest Pass district, 138,201 tons of 2,000 lb., that being the ton used throughout the Province. This production consisted of lignite, bituminous coal, anthracite, coking coal and coal used in making. briquettes, as follows: Tons "Lignite coal" (lignitic?) 1,763,225 Bituminous coal 2,374,401 Anthracite coal 168,720 Coking coal 104,012 Coke produced 65,167 Briquettes produced 130,861 The only district in which coke was made is the Crow&apos;s Nest Pass district, and the entire product was consumed in Canada. The total number of persons employed is as follows: Above ground 2,231 Underground 5,837 Total 8,068 The per capita production was as follows: Tons of coal mined per employee 533 Tons of coal mined per underground employee 737

Jan 7, 1914

By J. P. Heller, H. B. Bradley, A. S. Odeh

A potentiometric model technique is presented for determining the areal sweep efficiency of a five-spot well pattern, at and beyond breakthrough. A sharp interface between displaced and displacing fluids is assumed. Although the prototype svstem is referred to as a waterflood operation, obvious changes in the notation and computations will adapt the results to other displacement processes. The results of the study include the following. 1. Areal sweep efficiencies for a five-spot well pattern, at and beyond breakthrough, for mobility ratios (displacing to displaced fluid) of 4:1, 2:1, 1:1 and 1:4. 2. Extension of potentiometric analysis to the investigation of the areal sweep efficiency beyond breakthrough in a five-spot pattern by the application of conformal mapping and conductive-solid models. 3. The use of layers of conductive fabric in representing mobility ratio changes in potentiometric models. 4. The development of a probe mechanism for probing conductive solids. The results obtained by conductive-cloth models agree with earlier areal sweep efficiencies at breakthrough obtained by Aronofsky and Ramey 1 on the potentiometric analyzer using electrolytic-tank models. Results beyond breakthrough differ from those obtained by the X-Ray Shadowgraph technique. Data from this study show that, for mobility ratios greater than one, water cut rises rapidly as fluid is produced after breakthrough. However, for mobility ratios smaller than one, a large increase in area swept resulted with only a small increase in water cut. INTRODUCTION In calculating reservoir performance of waterflood-ing operations and other fluid-injection programs, it is necesssary to estimate the areal sweep efficiency * before and after injected-fluid breakthrough into production wells. Influence of mobility ratio on oil-production history, before and after breakthrough for a five-spot well pattern, has been studied by X-Ray Shadowgraph techniques 2-5 and by gelatin models.6 In none of these investigations was the transition zone controlled experimentally. Steady-state anaiog techniques assume that a vertical and discrete interface exists between the displacing and displaced phase. Thus, when using a potentiometric analysis, it is assumed that no transition zone exists. To determine the areal sweep efficiency beyond breakthrough for a five-spot well pattern using the potentiometric analyzer, four principal problems had to be resolved. These were the following. 1. The laborious method of representing changes in mobility ratio in electrolytic tanks by means of contoured wooden blocks. 2. Vapcrization of the electrolyte solution during the study causing concentration changes and, thereby, variable conductivity. 3. The scale-limitation problem 7 that is accentuated by abrupt eiectrolyte depth changes. 4. A wellbore geometry problem. EXPERIhlENTAL EQUIPMENT AND PROCEDURES CONDUCTIVE-SOLID MODEL The theory, equipment and procedural techniques of the electrolytic-tank potentiometric model have been treaced excellently by Lee8 and by Muskat.9 By the use of a conductive solid to represent the fluid conductivity of the porous media, the need for contouring wooden blocks to represent mobility ratio was eliminated. It was found that a carbon-black coated fabric, Uskon D-16,** could be employed as a fairly uniform conductive media. This material, when stacked in appropriate layers and pressed together, can be used to represent changes in mo-

By C. S. Finney, John Mitchell

The growing popularity in the United States of the vertical-flue even was emphasized when in 1905 the United States Steel Corp. chose the Koppers oven as the type which best suited their requirements. Heinrich Koppers was born on November 23. 1872. at a small farm in Walbeck near Geldern on the lower Rhine. When young Koppers was eight years old, however the family moved away from the farm to the industrial city of Bochum in the Ruhr. Here Koppers attended public school and subsequently served an apprenticeship to a tinsmith before taking a job as a lathe operator with a local steel company. He had ambitions to be much more than a machinist, however, and used his week-ends and evenings to improve his theoretical background by taking courses at a vocational-training school in Bochum. After winning the highest honor the school could bestow (the silver Staats-medaille), Koppers went on to continue his education at the Rheinisch-Westfalische Hüttenschule in Duisberg. One of his teachers there, Fritz Wüst who later became a professor at the Technische Hochschule at Aachen, recognizing Koppers&apos; unusual abilities, predicted for him a great future. In 1894 Heinrich Koppers joined the firm of Dr. C. Otto and Co. in Dahlhausen, and in 1899 while superintendent of the Mathias Stinnes mine he built his first battery of ovens for Hugo Stinnes, the German industrialist. Two years later he started his own organization, and in 1902 he made Essen his headquarters. It was to Essen that a group of engineers from the United States Steel Corp. went in 1906 with an invitation to Koppers to design and supervise the construction of four batteries of ovens at the Joliet works of the Illinois Steel Co. Each battery was to consist of 70 ovens. Arriving in the United States in 1907, Koppers established a branch of his firm in Joliet, and construction began. The first battery was fired on July 27, 1908. Rugged and simple, these ovens incorporated basic design features which were to make the Koppers oven and its future modifications the choice of a very large segment of the by-product coking industry of America. The 280 ovens at Joliet were 35 ft long, 8 ¾ ft in height, and tapered from 21 to 17 in. The total daily capacity of the four batteries was 2240 tons of coke. The ovens were of the new cross-regenerative type; that is, instead of longitudinal regenerators serving an entire battery, as in the older Koppers ovens, cross regenerators for each separate oven were employed. Fuel gas was supplied from the side of the battery through ducts in the brickwork known as gun flues, which reached to the center of the battery under the vertical heating-flues. Removable, ceramic gas-nozzles fitted at the top of each gun flue helped to insure good control over the distribution of the fuel gas, and uniform heating conditions were also promoted by regulating the air supply to, and the suction in, each heating flue. A different refractory w& used for each battery. One was built of American silica brick, one of American quartzite, and two of imported German quartzite. The installation at Joliet proved to be very successful, and in 1911, 490 additional&apos; Koppers ovens were built for the Illinois Steel Co. at the great new steelworks at Gary, Ind. By 1912 the H. Koppers Co. had established its headquarters in Chicago and was rapidly extending its business to include construction for such iron and steel companies as the Woodward Iron Co. at Woodward, Ma. (80 ovens in 1912); the Tennessee Coal, Iron and Rail- road Co. at Fairfield, Ala. (280 ovens in 1912); the Inland Steel Co. at Indiana Harbor, Ind. (86 ovens during 1913 and 1914) ; and the Republic Iron and Steel Co. at Youngstown, Ohio (68 ovens in 1913). In 1914 a group of men in Pittsburgh bought a major shareholding in the H. Koppers Co., and moved the headquarters of the organization from Chicago to their own city. Under its new management the company was highly successful in obtaining a large share of the contracts for by-product installations built during World War I. In 1917 the remaining German interests in the company were

Jan 1, 1961