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By Howard Pyle

CORE analysis is a recent development in the field of petroleum technology. The earliest work on this subject was done in connection with evaluating and planning secondary oil recovery by water-flooding. Present-day analysis of sands of flush fields requires new and independent interpretations of the data provided by cored material. Particular attention is being paid to the development of rapid, routine methods for measuring physical characteristics of sandstone, such as permeability, porosity and grain size, as well as the sand's fluid content. Permeability may be defined as the fluid-passing capacity of a rock. For a sandstone to be permeable it must be possible to pass a measurable quantity of fluid through the material in a finite period of time. The common unit of permeability is the darcy, which may be defined as the rate of viscous flow, in milliliters per second, of a fluid of one centipoise viscosity through a porous medium having a cross section of one square centimeter, under a pressure gradient of one atmosphere (76.0 cm. Hg) per centimeter.1 For convenience the subunit millidarcy, or one-thou-sandth darcy, is often used. Porosity is defined as the fluid-containing capacity of a rock and is usually expressed as a percentage ratio of the total pore volume to the rock or bulk volume. Since measurement of only the effective or communicating pore volume is often made, porosities are distinguished as total or effective.2,3 Net effective porosity is defined as the net pore volume available for oil and gas. Oil and water content of sand is measured and expressed either as the percentage saturation of the total pores or of the bulk rock. The oil content may also be expressed as the percentage saturation of the net effective pores. In connection with water-flooding, methods of core analysis were developed to furnish data that could be used in determining the quantity of recoverable oil,4-6 and the probable rates at which various sands or divisions of a sand open to the wells would take water or produce oil. In flush fields, the analytical data are used to predict the productivity of a cored interval. They indicate the presence of oil, gas and water and may

Jan 1, 1939

By Wm. R. Webster

This is a contribution for the proposed new discussion on the physics of steel. The former discussion on the subject started with the consideration of five papers presented at the Chicago meeting in 1893, and continued for several years. The Suggested Lines of Discussion, prepared by Doctor Howe, covered a wide field and added greatly to the success of the work. Naturally, the first step in a new discussion will be to remodel and bring up to date the old Suggested Lines of Discussion so as to embody all the important advances in the metallurgy of steel that have a bearing on the physics of steel and keep the papers and discussions within the field of work undertaken. This paper will be devoted to the practical application, in rolling steel, of the effects of carbon, phosphorus, and manganese on its tensile strength, with some suggestions on further research work. Dr. J. E. Stead, in his paper Influence of Some Elements on the Mechanical Properties of Steel,' gave a complete review of all that had been done during the past thirty years. He gave full credit to all investigators who worked on these lines, also their results, with his views on same, and in many cases his own results. All of this data and much new valuable information were put in convenient shape for reference and study. To do this required a great deal of work, which is fully appreciated by the steel manufacturers and investigators. The variations in the values given by the different investigators for the increase of the tensile strength of steel for each 0.01 per cent. of carbon, phosphorus, and manganese were largely due to the steels they worked on, the large or small variations in the amount of the elcments present, and the omission to take into consideration all the factors, from the blast furnace through to finished rolled material, that affect the character of the steel. For instance, a poorly made dirty steel requires considerably more carbon to give the same ultimate strength than a well-made clean steel with the same phosphorus and manganese content.

Jan 1, 1922

By Wm. R. Webster

This is a contribution for the proposed new discussion on the physics of steel. The former discussion on the subject started with the consideration of five papers presented at the Chicago meeting in 1893, and continued for several years. The Suggested Lines of Discussion, prepared by Doctor Howe, covered a wide field and added greatly to the success of the work. Naturally, the first step in a new discussion will be to remodel and bring up to date the old Suggested Lines of Discussion so as to embody all the important advances in the metallurgy of steel that have a bearing on the physics of steel and keep the papers and discussions within the field of work undertaken. This paper will be devoted to the practical application, in rolling steel, of the effects of carbon, phosphorus, and manganese on its tensile strength, with some suggestions on further research work. Dr. J. E. Stead, in his paper Influence of Some Elements on the Mechanical Properties of Steel,' gave a complete review of all that had been done during the past thirty years. He gave full credit to all investigators who worked on these lines, also their results, with his views on same, and in many cases his own results. All of this data and much new valuable information were put in convenient shape for reference and study. To do this required a great deal of work, which is fully appreciated by the steel manufacturers and investigators. The variations in the values given by the different investigators for the increase of the tensile strength of steel for each 0.01 per cent. of carbon, phosphorus, and manganese were largely due to the steels they worked on, the large or small variations in the amount of the elcments present, and the omission to take into consideration all the factors, from the blast furnace through to finished rolled material, that affect the character of the steel. For instance, a poorly made dirty steel requires considerably more carbon to give the same ultimate strength than a well-made clean steel with the same phosphorus and manganese content.

Jan 1, 1922

By E. P. Pfleider, D. H. Yardley

Layers of magnetic taconite of above average magnetic iron content extend down dip on the Mesabi Range to depths beyond the reach of open pit mining. A selected layer with thickness of 75 to 100 ft underlies many square miles. Such a selected layer can provide over 50 million tons of recoverable concentratc per square mile, at lower concentration ratios than in the present open pit taconite operations. The system contemplated would consist of large scale underground quarrying by a lane and pillar method. Working heights would be 60 to 100 ft; room widths, 50 to 75 ft; and productivity targets, about 40 tons per man-shift. The most critical problems would he roof control and blasthole drilling. The successful operation of large open pits and concentrating plants for treatment of magnetic taconites indicates that the basic techniques and economic problems of pellet production have been solved. The open pit reserves are large, but they are not necessarily available in sufficient quantity or quality to provide an adequate raw materials base for all potential producers. Furthermore, as the more easily available open pit crudes are used up, the pits will become deeper, stripping costs per ton of concentrate will increase, and recoveries will decrease as leaner crudes are utilized. The time will come when supplementary and replacement reserves of magnetic taconite will be required. One way these can be obtained is by underground mining of the down-dip extension of the Mesabi magnetic taconites, especially those layers averaging well over 25% recoverable iron. A key point in such a proposal is the possibility of realizing substantial savings by beneficiating an ore of 30% magnetic iron at 2.5 to 1 ratio of concentration. This contrasts with current practice in Mesabi commercial taconite plants, where the ore averages 22 to 25% magnetic Fe, with concentration ratios of about 3.0 when producing 61% Fe pellets. Two factors which favor the eventual development of underground taconite mining are: 1) the desirable physical structure and high grade of taconite pellets in general and 2) the improved engineering techniques in underground mining. Experience in large scale room and pillar mines demonstrates that, given strong rock strata immediately above a high grade mining

Jan 1, 1963

By P. W. George

This paper presents geological data regarding deposits of over 7,500,000 tons of fine-grained sulphide ore in barite gangue. A series of pyroclastics and arkoses was intruded by sills of quartz porphyry in connection with batholithic intrusions of granite 3 miles from Buchans. Gentle folding, with fracturing and shearing of tuff beds and intruded quartz porphyry, accompanied and followed the intrusion of granite and porphyry. The sulphide-barite mineralization is localized primarily in fractured and sheared tuff beds, but also in the sheared quartz porphyry. Two structures have proved particularly favorable for mineralization—one, the center of an anticline; the other, a series of weak, incompetent tuff beds intruded by thin sills of quartz porphyry and located between strong beds of arkose. During the folding the weak tuff with intruded quartz porphyry was intensely sheared and contorted and subsequently mineralized by sulphides and barite. Introduction Buchans is 5 miles north of Red Indian Lake, in Central Newfoundland, at an elevation of 900 ft. above sea level and 400 ft. above Red Indian Lake, and is reached by a 37-mile private railway from Miller-town Junction. The first ore body found at Buchans was discovered some 30 years ago in the bed of Buchans River, by Matty Mitchell, an Indian woodsman employed by the Anglo-Newfoundland Development Co. Ltd. In the years following the discovery the ore was developed by shaft-sinking and drifting, with results indicating a lenticular ore body in a shear zone, containing about 100,000 tons of extremely fine-grained sulphide ore in barite gangue averaging about 0.05 oz. Au, 4 oz. Ag, 1.5 per cent Cu, 10 per cent Pb, 18 per cent Zn. Because a profitable metallurgical treatment for the ore was not found, the mine was closed down in 1911. In 1915, Mr. H. A. Guess, Vice President of the American Smelting & Refining Co., learned of the Buchans River ore body, and after metallurgical tests had been carried on from time to time in his company's testing plant a satisfactory separation of the sulphides by selective flotation was finally effected in 1925. Prospecting for additional ore bodies was begun in 1926, utilizing geophysical methods under the direction of Hans Lund-

Jan 1, 1937

By Reinhardt Thiessen

Under sulfur in coal, is usually understood that form of sulfur which is combined with iron and known as pyrite. It occurs in the form of balls, lenses, nodules, continuous layers, thin sheets, or flakes, both in horizontal planes and vertical cleavage fissures. But pyrites also occur as very fine microscopic particles, or nodules, disseminated through the compact coal. This form has had but very little consideration. Finally, there is sulfur in coal in an amicroscopic form (not visible with an ordinary microscope), probably combined with the organic matter that exists in the coal. This form has had considerable attention from a scientific standpoint, but has probably not been recognized enough on the economic side. One or the other, or both, of these two latter forms may comprise the larger part of the sulfur content of coal, especially after it has been washed or otllerwise prepared for use. Pyrites in Microscopic Particles All coals that have been examined by the author contained a varying amount of sulfur in very small globules, or particles, of pyrite. These particles are seen in a thin [section as roughly rounded opaque dots, Figs. 1 to 10. When isolated, they are generally shown to be approximately spherical in shape, with a rough outer surface. They vary in diameter from a few microns to a hundred microns, the majority messuring from 25 to 40 microns; relatively few exceed the latter diameter. They are, therefore, very small objects. Their size is best appreciated by comparing the illustrations, in which they are shown. at a magnifiea tion of 100 diameters, with some known area like the period used in the ordinary printed page. The period is about 0.6 mm. in diameter, and . when magnified 100 times, will cover a circular area of about 2 1/2 in. in diameter, or all but the corners of the illustrations. The illustrations, therefore, represent an area about as large as an ordinary period, and the pyrite globules form but a small part of the total area. Very frequently a number of these particles are joined together in horizontal rows; oc-

Jan 1, 1920

By A. R. Rosenfield, G. T. Hahn

The effects of adding 0.025 volume fraction of uniformly dispersed Tho2, particles (particle size -380A) to iron have been examined. The two-phase alloy is shown to have a lower transition temperature and higher cleavage stress than either the base metal or a solid-solution alloy (Fe-3Si). Enhanced cleavage resistance is thought to be a consequence of particles limiting the effective slip distance of dislocations and thus the ability of Pile-ups to act as stress concentrations. Further calculations predict that the highest cleavage strengths are attained with uniform dispersions, where the particle size is kept small (-250A) and the volume fraction is relatively large (-2 to 3 pct). ALTHOUGH a fine dispersion of second-phase particles is known to impart strength to an alloy, the possibility that it might also impart ductility has only recently been explored. Outstanding progress has been made in the production of ductile tungsten alloys containing Tho21,2or (Hf, Zr)C,= but the contribution of the dispersion to ductility has neither been optimized nor fully analyzed. The basis for such an analysis has now been developed1,5 and is discussed briefly below. Experiments on iron are also described that attempt to isolate one of the possible ductilizing mechanisms and test its effectiveness. In analyzing these experiments, it is important to remember that cleavage obeys a stress criterion, and that either a lower ductile-to-brittle transition temperature or a higher cleavage stress is evidence of enhanced resistance to cleavage. DUCTILIZING MECHANISMS Slip Blockage. Cleavage fracture of polycrystals may be initiated at the intersection of a slip band (or twin) and a neighboring grain or a large (-1 to 5 µ) particle. The slip band generates a stress concentration at its head, and either the grain boundary fractures, the particle fractures, or cleavage is initiated in a neighboring grain. Experimental evidence for this viewpoint is found in metallographic studies8-8 and from the observation that cleavage is usually preceded by yielding.8-11 The potency of a slip band as a stress concentrator

Jan 1, 1968

By John T. Crawford

Much has been written of late regarding the flotation of coal as a means of reclaiming the valuable portion of the fines encountered in nearly all methods of coal preparation. Whether the process be wet or dry, whether the dust be sent through the entire process or removed prior to the cleaning by some such method as screening or aspiration, one of the final products consists of a considerable proportion of — 48-mesh material. The percentage of this —48-mesh material depends upon the mining method, transportation, storage, and method of preparation. The tonnage of this dust material ranges from 1 to 10 per cent of the total output of the cleaning plants. Four methods of disposal are available for this coal: 1. Disposal with the refuse from the plant. 2. Addition, after being reclaimed from the wash water, or dust collectors, without cleaning to the larger sizes. 3. Separate disposal as a pulverized fuel. 4. Cleaning so that addition to the larger sizes benefits the product as a whole. It is also true that flotation in one form or another is the only known way to effectively clean this material. While some processes do a reasonable cleaning job down to and including +l00 mesh, by far the greater majority stop at or short of the + 48-mesh size. Much of the literature on coal flotation passes by the fred pulp density and size by merely stating that the feed varies from 20 to 30 per cent solids and that the size range is from — 1/4 in. or — 1/8 in. to 0. Small wonder then that some operators with entirely different conditions, and not knowing the machine capacity, throw up their hands in despair when flotation is mentioned. Too much parallel is drawn between flotation methods as applied to ores and as applied to coal, to the detriment of the latter. In the flotation of metalliferous ores the feed pulp density ranges from 20 to 30 per cent solids, depending upon the amount of mineral to be floated and upon the size analysis. Thus where 2 to 5 per cent of the

Jan 1, 1936

By John T. Crawford

Much has been written of late regarding the flotation of coal as a means of reclaiming the valuable portion of the fines encountered in nearly all methods of coal preparation. Whether the process be wet or dry, whether the dust be sent through the entire process or removed prior to the cleaning by some such method as screening or aspiration, one of the final products consists of a considerable proportion of — 48-mesh material. The percentage of this —48-mesh material depends upon the mining method, transportation, storage, and method of preparation. The tonnage of this dust material ranges from 1 to 10 per cent of the total output of the cleaning plants. Four methods of disposal are available for this coal: 1. Disposal with the refuse from the plant. 2. Addition, after being reclaimed from the wash water, or dust collectors, without cleaning to the larger sizes. 3. Separate disposal as a pulverized fuel. 4. Cleaning so that addition to the larger sizes benefits the product as a whole. It is also true that flotation in one form or another is the only known way to effectively clean this material. While some processes do a reasonable cleaning job down to and including +l00 mesh, by far the greater majority stop at or short of the + 48-mesh size. Much of the literature on coal flotation passes by the fred pulp density and size by merely stating that the feed varies from 20 to 30 per cent solids and that the size range is from — 1/4 in. or — 1/8 in. to 0. Small wonder then that some operators with entirely different conditions, and not knowing the machine capacity, throw up their hands in despair when flotation is mentioned. Too much parallel is drawn between flotation methods as applied to ores and as applied to coal, to the detriment of the latter. In the flotation of metalliferous ores the feed pulp density ranges from 20 to 30 per cent solids, depending upon the amount of mineral to be floated and upon the size analysis. Thus where 2 to 5 per cent of the

Jan 1, 1936

By C. D. Woodward

PRIOR to 1902, the tramming of ore from the stopes to the shafts, in the Butte mines, was done by man or animal power, but the demand for greater tonnage and the need for more improved methods of tramming the ore resulted in the installation of the compressed-air locomotive in a few mines. Some of the objections to animal haulage were thus eliminated, but this locomotive was not generally adopted, because of a number of limiting features. It was superseded by the electric trolley locomotive. The early types of trolley locomotives were not satisfactory from the operators' viewpoint, but. they had many economic advantages and after a number of years, in which improvements were constantly made in the structural details, they became dependable, so that the use of animals for main haulage in the mines has now been practically abandoned. There are now operating in the Butte mines of the Anaconda Copper Mining Co., 185 trolley-type locomotives, each of which has replaced five animals. It cost approximately $50 per month to keep an animal, hence the economies that resulted fully justified the expenditure of electrifying the haulage systems. The haulage conditions in this company's mines possess many advantageous features that are not always found in underground mine workings. The drifts and crosscuts are large, approximately 5 ft. (1.5 m.) wide by 7 ½ ft. (2.3 m.) high, and have curves of no less than 18 ft. (5.5 m.) radius; and the roadbeds of the drifts and crosscuts have a uniform grade of 0.5 per cent. toward the shaft, or in favor of the loaded trips. In all of the mines, except two, the track gage is 18 in. (45.7 cm.), and all tracks in the main haulageways are laid with 25-lb. (11.3 kg.) rails. The stations at the shaft are large, well-lighted, clean, and cool. The average temperature of ,the stations, however, is about 75° F. (24° C.), while the average temperature in some of the drifts and crosscuts is 92° F. (33.5° C.).

Jan 2, 1922

By William L. Fink, H. R. Freche

The constitution of the aluminum-cobalt alloys has been the subject of relatively few investigations. Brunckl analyzed the residue which remained after treating the 17 per cent cobalt alloy with dilute hydrochloric acid. The results of the analyses indicated the existence of a compound Co3Al13. Guillet2 published a liquidus for the aluminum-cobalt system which contained two maxima corresponding to CoA16 and CoAl2 and two eutectic points at 30 and 74 per cent Co. The most extensive investigation of the system was carried out by Gwyer13 whose results are summarized in Fig. 1. It will be noted that at the aluminum end there is a suppressed eutectic at 644' C. By extrapolation of the duration of the eutectic arrest to zero, the composition of the compound was placed at Co3Al13. Another thermal arrest was noted at 550" C., which Gwyer attributed to a peritectic reaction. The present investigation of the aluminum end of the aluminum-cobalt system is the eleventh of a series dealing with the constitution of aluminum-rich alloys prepared from electrolytically refined a1uminum.l Preparation of Alloys.—The alloys used in this work were all prepared from electrolytically refined aluminum, the analysis of which is shown in Table 1. In preparing some of the alloys, metallic cobalt (98 to 99 per cent) was used either directly or in the form of an intermediate alloy, the composition of which is given in Table 1. In order to obtain alloys of higher purity, cobalt was introduced into some of the alloys by means of chemically pure cobaltous chloride which was readily reduced by the aluminum. A small electric furnace with nickel-chromium heating elements was used in the preparation of the alloys. Chemical Analysis.—An analysis sample from each melt was chill-cast in the form of a plate 3/16 in. thick. In all cases the casting temperature was well above the liquidus. Cobalt was determined electrolytically in all alloys containing more than 0.05 per cent cobalt and by the nitroso-

Jan 1, 1932

By S. K. Nowak

The lithium solubility limit in solid aluminum was determined by the use of micro-graphic techniques. The solubility limit thus established was shown to be a true equilibrium by checking the reversibility of AlLi phase precipitation and dissolution around the equilibrium temperature. The experimental results fit a straight line by plotting log N(LI) against 1/T. The metallurgical reasons explaining the lack of agreement among previous investigations are discussed. THE solubility limit of lithium in solid aluminum was determined by Shamray and Saldau1 by micrography and by Grube et al.' and also Voss-Kiihler- by measurement of electrical resistivity. Since the addition of lithium induces extremely small changes in the aluminum lattice, X-ray techniques could not be employed for this purpose; in fact, the addition of lithium does not cause an expansion of the lattice as might be expected from a consideration of atomic radii, but causes a small contraction from 4.040 to 4.037A3, 4 when the solvent aluminum is saturated with this element. An analysis of the results from previous investiga-tions'" shows that with improvement of investigation techniques the solubility limit tends toward a lower lithium content. Again, some of the author's own experimental results on the action of atomic hydrogen on A1-Li alloys seemed to indicate that the solubility of lithium should be much lower, i.e., about 0.3 pet at 300°C: instead of 1.15 pet, the lowest published value.4 For the foregoing reasons it was decided to investigate this problem. The aluminum and the lithium used for this purpose were of the highest purity. The lithium solubility limit was determined by metallographic techniques, in particular by the use of a reagent especially developed for the purpose of revealing the AlLi phase. The solubility limit thus determined was shown to be a true equilibrium by checking the reversibility of AlLi phase dissolution and precipitation around the equilibrium temperature. All of the experimental work was carried out in 1951 at the Centre d'Etudes de Chimie Metallur-gique at Vitry, France, under the direction of Georges Chaudron.

Jan 1, 1957

By E. P. Pfleider, O. Rellensmann

For many years mine surveyors and exploration engineers have sought an accurate means of transferring meridian underground by using the gyro-compass. These efforts have generally failed, either because they were inaccurate or because the size and cost of the unit were prohibitive. During the past ten years, however, great strides have been made. Autonetics in California has developed a comparatively light-weight (100-lb) gyro-theodolite, and the Institute for Mine Surveying at Clausthal Mining Academy in Germany has developed an instrument that is accurate to within 20" to 30" of arc. Concurrently, Minneapolis-Honeywell has produced an astounding gyro, the size of a salt shaker, that weighs less than 0.5 lb and compares favorably in performance with the 50-lb gyros of conventional design. This unit, designed for missile guidance, should be adaptable for surveying small boreholes. Given the proper interest and approach from the engineering fraternity, excellent gyrocompasses could probably be available for accurate underground survey work at nominal cost within the next three years. Principles of the Gyro-Compass: The gyrocompass, used for 50 years as a navigational instrument in ships and aircraft, senses the turning axis of the earth and points out true polar north. It does this by employing the earth's rotation, the force of gravity, and the precession characteristics of a gyro. For further details the reader is referred to several excellent publications'"3 covering the principles and application of the gyroscope. The heart of the gyro-compass is the gyroscope, or rapidly spinning wheel, which rigidly maintains its angular orientation in inertial space. Free gyros employ universal joints to allow the wheel axle to maintain its rigid spatial orientation independent of base motion disturbances. Free gyros are independent of the earth's turning and are not north-seeking. Only as they might be initially oriented parallel to the earth's turning axis could any north indication be obtained. In practice, the free gyro tends to drift off its set position and lose its accuracy. The Sperry-Sun Well Surveying Co. instrument (5M in O. D.) is of this free gyro type. It is counterbalanced, however, against the rotational effects of the earth, so that orientation of the spinning axis will be maintained throughout the survey.

Jan 1, 1960

By F. D. Rosi

IN the production of single crystals by the Bridg-man method of solidification from the melt in vacuum at a crucible lowering rate of 0.25 in. per hr, a cellular structure was frequently observed in copper and copper-base alloys at the terminal surface of the as-cast crystals (i.e., the surface representing the last portion of the crystal to solidify). The geometrical form of this structure varied from crystal to crystal, but, in general, could be classified as either hexagonal or quadrilateral. Fig. 1 shows examples of this cellular structure, which is similar to that reported by Buergerl and, more recently, by Pond and Kessler&apos; and Rutter and Chalmers. In the micrographs of Figs. lb and c, the cubic structure in the alloy crystals is obscured somewhat by excessive surface evaporation in the final stages of solidification. X-ray orientation studies indicate that the traces outlining the various geometrical forms correspond to octahedral planes, which are the planes of greatest reticular density in the face-centered cubic lattice. It would appear from this that the view first propounded by Bravais,4 that the surface energies and normal growth velocities of the different crystal planes are inversely proportional to their reticular densities, applies to this cellular structure, since those surviving are the planes of greatest atomic density. Thus, the observed variation in the geometrical form of this structure in crystals of different orientation is readily understandable from the distribution of the (111) planes in the standard {111}, (1101, and (1001 stereographic projections. Depending upon whether the crystal axis (i.e., the direction of crystallization) coincides with a <111>, <111>, or <100> direction, the cellular structure viewed from a surface perpendicular to the crystal axis will consist of a network of regular hexagons, skewed hexagons (with interfacial angles of 110" and 125"), or cubes, respectively. It follows, moreover, that off-orientations from these principal directions will yield modifications of these three basic geometric configurations. In this connection, it is interesting to note that the bounding planes of the hexagonal structure observed in tin by Rutter and Chalmers3 showed no preferred orientation.

Jan 1, 1954

By William A. Edmiston, R. E. Grace

A simplified model for the theory of reaction rates near equilibrium has been applied to the reduction of wustite. A linear relationship between the net rate of reaction and the Gibbs free-energy change, ?F, is predicted. The reduction of wustite in H2 -H2O gas mixtures was studied with a microbalance technique at temperatures from 850° to 950 °C. The predicted model was found to be valid to -?F/RT = 0.1 or approximately -200 cal per mole away from equilibrium. ThE theory of reaction rates near equilibrium has been developed by a number of authors. Prigogine, Outer, and Herbo1 have studied the catalytic hydrogena tion of benzene in an attempt to deduce some intrinsic connection between reaction rates near equilibrium and thermodynamics. Manes, Hofer, and weller2 have derived more general expressions for reaction rates near equilibrium from classical thermodynamics. Hollingsworth3,4 has commented briefly on the work of Manes et al. Benson5 has summarized in a general way the results of the earlier contributions. It is the purpose of this paper to apply the theory of reaction rates near equilibrium specifically to the reduction of wustite in H2-H2O gas mixtures. The basic premise is that the reduction of wustite proceeds by a surface reaction, rather than by solid-state diffusion ,or gaseous diffusion. Richardson and Dancy6 and Gellner and Richardson7 have suggested that the reduction of wustite may be governed by iron diffusion in wustite. Mckewan8-11 has proposed that reduced iron layers on various oxides are sufficiently porous to offer a negligible impedance to the reducing gas. The linear motion of interfaces reported by McKewan and discussed by Themelis and Gauvin12 indicates the lack of diffusive control during reduction. The reduction studies of magnetite reported by Quets, Wadsworth, and Lewis13 give additional evidence that reduction rates are linear when the reaction products are porous and nonprotective. The interpretation of measurements reported in this paper is predicated upon a surface-controlled reduction reaction, at least in the earliest stages of reduction when the reduced iron layer is thin.

Jan 1, 1967

By R. G. Knickerbocker, W. H. Yarroll, P. E. Churchward, J. W. Hunter, J. H. Jacobs

THE present paper records a chapter in the history of the development of an electrolytic manganese industry in the United States.1 A relatively large pilot plant at Boulder City, Nev., for the production of electrolytic manganese was conceived as part of the national defense program to make this country independent of foreign importation of manganese ore. The first authorization, in November 1940, however, was for a plant of only several hundred pounds capacity. Improvements in process and subsequent additions to some sections authorized by subsequent appropriations have enabled the plant to produce as much as 2000 lb. of manganese a day for substantial periods of time. Since in pilot-plant operation frequent changes are made in the flowsheet, the electrolytic manganese plant was constructed so that equipment could be rearranged with a minimum of expense and time. Several different kinds of the various types of equipment were installed for many of the operations, so that the one most suitable for the process might be chosen. Pilot-plant operations were begun in November 1941 and have been continuous except for brief shutdowns for installation of new equipment. More than one-half million pounds of manganese have been produced. The improvements in process, selection of equipment, and operational procedure developed are recorded here. DEVELOPMENT OF THE PROCESS The work of early investigators on the electrolysis of manganese solutions has been reviewed by Shelton.2 Until the method of continuous production of manganese by electrolysis was perfected by Shelton2-4 and coworkers, no practical procedure had been devised. The method as originally developed by the Bureau of Mines will be described here only briefly. The manganese dioxide ore was reduced in a rotating kiln-type furnace at 650°C. in an atmosphere of "city" gas. The calcine was leached with spent electrolyte and enough sulphuric acid to produce a neutral leach solution containing 25 grams manganese per liter as manganese sulphate and 200 grams per liter of ammonium sulphate. Arsenic was removed from the solution by the addition of 0.05 gram iron per liter as ferrous sulphate followed by oxidation

Jan 1, 1944

By Eric J. Jette

In thermodynamic studies of gas mixtures and liquid solutions, the respective problems have been greatly simplified by the use of two general limiting laws; Dalton&apos;s law of partial pressures and Raoult&apos;s law. It is generally recognized that these two laws have a common physical and chemical basis which may be stated as follows: The law may be expected to be strictly valid only for systems in which the different molecular species are so nearly the same in their physical and chemical properties that a molecule is unable to recognize whether its neighbor is one of its own kind or not. Positive and negative deviations from both laws are frequently encountered. The occurrence of negative deviations in both gases and liquids is ascribed to attractive forces which, in the more extreme cases, may be so large in magnitude and so highly specific in character as to lead to the formation of definite chemical compounds. Positive deviations in gases develop only at higher pressures where the intermolecular distances are so decreased that factors such as molecular size and shape, which ultimately become concerned with the repulsive forces, come into play. In liquid mixtures, positive deviations are generally considered as due to differences in the symmetrical (nonpolar) fields of force which surround the various molecular species and which are supposed to give rise to a tendency for each species to associate with its own kind and thus to "squeeze out" the other variety. Fundamentally this also involves the existence of repulsive forces between unlike molecules. If the fields of force are unsymmetrical (polar) a tendency towards compound formation results. Stated in more exact terms, strict adherence to the law means that the "activity" (or any adequate measure of this quantity) of a component in a mixture is a linear function of the molecular (or atomic) fraction of that component between the mol fraction limits 0 and 1. A positive deviation means that the activity is greater than corresponds to the law; a negative deviation that the activity is less. The deviations from Raoult&apos;s law have been fully discussed by many writers1

Jan 1, 1934

By J. W. R. Rayner

THE conventional method of making water gas involves individual plants for the separate carbonization of coal to coke and the subsequent gasification of coke with steam. The process demands lump coke, preferably made from the best quality of coking coals. The ash in the coke should have a high melting point and the gasification process suffers a loss of efficiency when cokes with law ash-melting points are used. In Britain, these coals are generally the most expensive and they are readily available only in certain localities. The coals not suitable for making lump coke, or yielding only poor quality coke, are much more widely available and are generally cheaper. DEVELOPMENT OF ALTERNATIVE PROCESSES In Germany, processes have been developed and operated commercially for the production of water gas from very low-rank coals such as brown coal and lignites, but much less effort has been spent on processes to use coals of intermediate rank, which are widely available in Britain. The brown-coal processes use coal in the form of either dust or briquettes. The most successful of the dust-gasification methods is the Winkler process, which made a substantial contribution to German oil production during the war. It uses coal in the site range 0 to 8 mm and the fuel bed is blown with a mixture of oxygen and steam at a rate high enough to give it a characteristic boiling motion or, in present-day phraseology, at a rate more than sufficient to fluidize it (Fig. 1). Air cannot be used to provide the heat required for gasification, for various reasons. In the first place, to make the process continuous by mixing air and steam would involve the production of a lean gas-containing much nitrogen. If a separate blow were used, a large amount of potential heat would be lost because the blow gas in such a process would contain a very high percentage of carbon monoxide. Oxygen is necessary therefore and oxygen remains relatively expensive for use in gasification. In the Winkler generator, formation of clinker caused by local overheating does not occur, because of the mobility and high heat transfer of a fluidized bed. These properties of fluidized dusts have been used most successfully

Jan 1, 1953

By C. D. Woodward

PrioR to 1902, the tramming of ore from the stopes to the shafts, in the Butte mines, was done by man or animal power, but the demand for greater tonnage and the need for more improved methods of tramming the ore resulted in the installation of the compressed-air locomotive in a few mines. Some of the objections to animal haulage were thus eliminated, but this locomotive was not generally adopted, because of a number of limiting features. It was superseded by the electric trolley locomotive. The early types of trolley locomotives were not satisfactory from the operators&apos; viewpoint, but they had many economic advantages and after a number of years, in which improvements were constantly made in the structural details, they became dependable, so that the use of animals for main haulage in the mines has now been practically abandoned. There are now operating in the Butte mines of the Anaconda Copper Mining Co., 185 trolley-type locomotives, each of which has replaced five animals. It cost approximately $50 per month to keep an animal, hence the economies that resulted fully justified the expenditure of electrifying the haulage systems. The haulage conditions in this company&apos;s mines possess many advantageous features that arc not always found in underground mine workings. The drifts and crosscuts are large, approximately 5 ft. (1.5 m.) wide by 71/2 ft. (2.3 m.) high, and have curves of no less than 18 ft. (5.5 m.) radius; and the roadbeds of the drifts and crosscuts have a uniform grade of 0.5 per cent. toward the shaft, or in favor of the loaded trips. In all of the mines, except two, the track gage is 18 in. (45.7 cm.): and all tracks in the main haulageways are laid with 25-lb. (11.3 kg.) rails. The stations at the shaft are large, well-lighted, clean, and cool. The average temperature of the stations, however, is about 75" F. (24" C.), while the average temperature in some of the drifts and crosscuts is 92" I?. (33.5" C.). As the orebodies lie at varying distances from the hoisting shafts,

Jan 1, 1923

By C. D. Woodward

PrioR to 1902, the tramming of ore from the stopes to the shafts, in the Butte mines, was done by man or animal power, but the demand for greater tonnage and the need for more improved methods of tramming the ore resulted in the installation of the compressed-air locomotive in a few mines. Some of the objections to animal haulage were thus eliminated, but this locomotive was not generally adopted, because of a number of limiting features. It was superseded by the electric trolley locomotive. The early types of trolley locomotives were not satisfactory from the operators&apos; viewpoint, but they had many economic advantages and after a number of years, in which improvements were constantly made in the structural details, they became dependable, so that the use of animals for main haulage in the mines has now been practically abandoned. There are now operating in the Butte mines of the Anaconda Copper Mining Co., 185 trolley-type locomotives, each of which has replaced five animals. It cost approximately $50 per month to keep an animal, hence the economies that resulted fully justified the expenditure of electrifying the haulage systems. The haulage conditions in this company&apos;s mines possess many advantageous features that arc not always found in underground mine workings. The drifts and crosscuts are large, approximately 5 ft. (1.5 m.) wide by 71/2 ft. (2.3 m.) high, and have curves of no less than 18 ft. (5.5 m.) radius; and the roadbeds of the drifts and crosscuts have a uniform grade of 0.5 per cent. toward the shaft, or in favor of the loaded trips. In all of the mines, except two, the track gage is 18 in. (45.7 cm.): and all tracks in the main haulageways are laid with 25-lb. (11.3 kg.) rails. The stations at the shaft are large, well-lighted, clean, and cool. The average temperature of the stations, however, is about 75" F. (24" C.), while the average temperature in some of the drifts and crosscuts is 92" I?. (33.5" C.). As the orebodies lie at varying distances from the hoisting shafts,

Jan 1, 1923