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By D. O. Ehrenburg, R. J. Watson

During the earlier period of electrical prospecting, the search for orebodies was by far the most important application of this method of geophysical prospecting. In the past few years, however, increasing attention has been paid to the possibilities of this method in the investigation of subsurface structures. Not only have fault zones been located, but attempts have been made to determine the depth to certain strata and thus contouring on some bed which proves to be a good "electrical marker." Many workers have published the results of field tests showing how their particular interpretations have agreed with the actual field conditions, but very few have stated their method of interpretation. Weaver, Hummel and Stefanesco and Schlumbergerl have given a more theoretical discussion of the resistivity methods. Hummel and Stefan-esco and Schlumberger have published mathematical solutions for the flow of electricity in a medium involving four layers. In this paper the authors present a solution for the flow of electricity in a medium of any number of horizontal strata. By means of this solution, curves showing the relationship between the apparent resistivity and the electrode spacing can be computed for ideal cases. Point-source in an Infinite Homogeneous Medium In resistivity measurements the current is conveyed to the earth by means of two electrodes, which can be mathematically treated as point-sources of equal and opposite strength. A point-source of strength I is defined as a point at which electricity is being generated at the constant rate of I coulombs per second, or I

Jan 1, 1932

By Hallan N. Marsh

The subject of mud sounds so simple, uninteresting and unimportant that it has failed to receive the attention that it deserves, at least as applied to the drilling of oil wells. As a matter of fact, it is one of the most complicated, technical, important and interesting subjects in connection with rotary drilling. In 1923, R. E. Colloml said "It should be possible to establish certain physical standards for mud fluid," but while considerable work has probably been done on the subject before and since that time, and numerous articles have been written, the progress of general knowledge of the subject and the application of better practices have not kept pace with other developments in rotary drilling, and physical standards cannot be considered as established. Mud is apt to be thought of as an accessible and cheap material, but its cost is likely to be an important factor in the total cost of drilling a well. Some wells "make" most of their own mud out of the formations drilled, but in most cases a great deal of mud-making material has to be supplied from more or less distant sources. Expenditures for mud materials by one California company totaled practically one-quarter million dollars during 1929 and averaged about $13 per rig per drilling day, or about $2000 per well. Under unfavorable conditions, mud cost may reach many times this figure. Parker2 says that in the Hobbs, N. M., field "as much as $75,000 has been spent on one well alone for mud and mud-weighting compounds."' Presumably this is an exceptional case. Unit cost of mud fluid will range from 10 c. per barrel, where suitable material is close at hand, to $1 where crude material is available at moderate price but must be hauled long distances, and on up to as much as $11 per barrel where high gas pressure requires the use of special processed weighting material. The cost of mud pales into insignificance, however, when the results of using improper mud are considered. Blow-outs and lost circulation are obviously the result of inadequate mud. Stuck drill pipe, whether or not twisted off, in many if not most cases probably is caused by cuttings settling out of the mud or the mud itself precipitating. Recent experi-

Jan 1, 1931

By Donald A. Cumming

Since the first major mine development, the company has continually looked for ways of improving drifting methods. The planned increase in production to 60,000 tpd, and the lack of trained personnel, requires new development methods to maintain this high production rate. In order to finish the drifting on the 2015 and 2075 levels, a total of 59,000 ft is required, or two years at the present rate.

Jan 4, 1972

By Paul J. Bennett, James E. Castle

The kyanite minerals, including kyanite, silimanite, and andalusite, are anhydrous aluminum silicates with the formula Al2O3SiO2. lumortierite and topaz are also included in this group because they are closely allied in composition and thermal behavior. However, leither are mined commercially today. They ire typical metamorphic minerals which are found in metamorphic rocks on every continent. Kyanite minerals are prized chiefly for their refractoriness and are important components in a broad range of acid refractory products, especially in mortars and castables. World production of kyanite minerals is currently about 250,000 tpy and has been increasing at the rate of 3 to 8% a year for many {ears (Anon, 1972). While kyanite minerals lave widespread occurrence, the consumption of kyanite minerals is concentrated in the relatively few highly industrialized areas where refractories are manufactured, and which in turn are typically close to the major iron and steel producing regions of the world. Thus northern Europe, England, the United States, and Japan are the principal consumers of refractories and kyanite minerals. Of these countries only the U.S. is a significant producer. The largest use for domestic kyanite, both raw and calcined, is in the manufacture of refractory mortars, cements, castables, and plastic ramming mixes. In these applications kyanite constitutes from 10 to 40% of the mixture, the balance being refractory clays and coarser grog materials. A certain proportion of raw kyanite is used to offset the shrinking of the clay binder, whereas the calcined kyanite is used in the coarser sizes for body. Certain high grade refractory shapes and kiln furniture are composed chiefly of graded sizes of calcined kyanite bonded with a little ball clay. In most applications, however, kyanite is used only to fortify the mix or as a minor ingredient. Some manufacturers of mullite brick use domestic calcined kyanite for the fine-grained portion of the brick, but seldom in quantities exceeding 10% by weight. Kyanite is used by ceramic manufacturers of wall tile and sanitary ware to offset shrinkage and cracking after firing. Indian kyanite has traditionally been used to make the coarse grog sizes needed for the manufacture of mullite brick, and for the grog portions of mortar, cement, and castable mixes. It has been largely replaced, however, by synthetic mullite, high fired flint, and diaspore clay or calcined bauxite for other applications. Indian kyanite is available in lump as well as in granules and concentrates, as are sillimanite materials from India and South Africa. Andalusite is available as 8 mesh granules and as fines with chemical analyses not unlike those of domestic kyanites. From time to time and from place to place minor amounts of other aluminum silicate minerals such as dumortier- ite, topaz, and pyrophyllite have been made available at various prices and specifications, but have not been significant factors in the refractory industry. A mixture of kyanite and sillimanite of relatively fine grain-size and of low iron content was at one time offered to the molding sand industry by the Du Pont Co. as a byproduct from their deposits of heavy

Jan 1, 1975

AN excerpt from the Odessa Isvestia Vechernia, dated April 24, 1927, and translated by P. Routzitzky, summarizes a report by Professor Preobragensky ad-dressed to the Russian Government. Professor Preo-bragensky writes in regard to newly discovered potash deposits in eastern European Russia. Near Solikamsk in the Perm district, about 100 miles east of the Ural mountains and on the Kama river, large deposits have been located and the belt has been traced southward about 25 miles. Subsequently the search for potash salts has been transferred to the Berezovsky district, near Nikolsk, on the Kosva river, about 70 miles north-east-of the city of Perm. Toward the end of March, 1927, at a depth of 800 ft., drilling at the Levensk salt works revealed potash salts in, definite strata. The bed has a thickness which extends from a depth of- 245 to 307 meters (latest data in middle of April, 1927) and has not been bottomed yet. Professor Preobragensky goes on to say that these operations have definitely, developed very large de-posits, the limits of which are still unknown. There is at least an implication that the reserves far exceed those of Stassfurt, Germany. For European use they are, of course, much less accessible, for by through rail they lie 1200 miles east of Petrograd. Further, no particulars are given as to the potash content. It is to be supposed that the salt occurs in the Permian beds that extend north and south along the west flank of the Ural mountains, as the localities mentioned .all lie within the area so mapped on the new geologic map of Russia issued by the Scientific and Technical Di-vision of the All-Russian Council of Soviets in 1922. It is interesting to note that these salt beds are of the same geologic age as those of 'the Stassfurt region and of the west Texas-New Mexico district, in which the U. S. Bureau of Mines is at present drilling.

Jan 7, 1927

By David B. Reger

Sharp declines of initial production of both oil and gas occurred in West Virginia during 1943. There was only a small decrease in the number of ncw wells, but these wells did not show the quantities of oil and gas that might have been expected. The same situation prevailed both as to routine drilling in old pools and as to exploratory wildcat wells. Out of 42 wildcats drilled, 19 resulted in small commercial gas wells, and 2 in small commercial oil. No new oil pools were discovered, but the proved limits of several old pools were considerably extended, giving 4200 acres of newly proved oil territory. Three new gas pools, resulting from discoveries made in 1942 and former years, came into definite activity. The newly proved gas territory in these and in various extensions of old pools is 39,000 acres or more. It is doubtful whether the newly proved reserves of either oil or gas will balance current withdrawals. The account of operations, as gathered from trade journals and other reporting services, shows that 696 new wells were drilled, resulting in 82 oil wells with 919 bbl. of daily new production; 475 gas wells with 454,735,000 cu. ft. of daily open flow, and 137 dry holes. Also, 57 old wells were drilled to deeper sands, with 3 bbl. and 11,484,000 cu. it. of added production. On the new wells, the oil average was 11.2 bbl. per well per day, and the gas average was 957,337 cu. ft. per well per day. On new wells the ratio of dry holes to completions was 19.68 per cent. On deeper drilling the ratio of failures was 14.03 per cent. The greatly decreased average initial gas-well production in 1943, as compared with 2,270,406 CU. ft. per well in 1942, is directly attributable to the abrupt decline of completions in the Charleston gas field, which reached its peak in 1942. Production of oil for the year was estimated by the U. S. Rureau of Mines as 3,349,000 bbl., as compared with 3,574,000 bbl. in 1942. Production of natural gas for the year is estimated by the author as 210,000, 000,000 cu. ft., compared with the U. S. Bureau of Mines final figures of 215,193,-000,000 cu. ft.* in 1942.

Jan 1, 1944

By David B. Reger

Sharp declines of initial production of both oil and gas occurred in West Virginia during 1943. There was only a small decrease in the number of ncw wells, but these wells did not show the quantities of oil and gas that might have been expected. The same situation prevailed both as to routine drilling in old pools and as to exploratory wildcat wells. Out of 42 wildcats drilled, 19 resulted in small commercial gas wells, and 2 in small commercial oil. No new oil pools were discovered, but the proved limits of several old pools were considerably extended, giving 4200 acres of newly proved oil territory. Three new gas pools, resulting from discoveries made in 1942 and former years, came into definite activity. The newly proved gas territory in these and in various extensions of old pools is 39,000 acres or more. It is doubtful whether the newly proved reserves of either oil or gas will balance current withdrawals. The account of operations, as gathered from trade journals and other reporting services, shows that 696 new wells were drilled, resulting in 82 oil wells with 919 bbl. of daily new production; 475 gas wells with 454,735,000 cu. ft. of daily open flow, and 137 dry holes. Also, 57 old wells were drilled to deeper sands, with 3 bbl. and 11,484,000 cu. it. of added production. On the new wells, the oil average was 11.2 bbl. per well per day, and the gas average was 957,337 cu. ft. per well per day. On new wells the ratio of dry holes to completions was 19.68 per cent. On deeper drilling the ratio of failures was 14.03 per cent. The greatly decreased average initial gas-well production in 1943, as compared with 2,270,406 CU. ft. per well in 1942, is directly attributable to the abrupt decline of completions in the Charleston gas field, which reached its peak in 1942. Production of oil for the year was estimated by the U. S. Rureau of Mines as 3,349,000 bbl., as compared with 3,574,000 bbl. in 1942. Production of natural gas for the year is estimated by the author as 210,000, 000,000 cu. ft., compared with the U. S. Bureau of Mines final figures of 215,193,-000,000 cu. ft.* in 1942.

Jan 1, 1944

By H. S. Price, D. A. T. Donohue

The system of equations describing displacement of a hydrocarbon liquid by a hydrocarbon vapor in a porous medium where mass transfer takes place between the phases is solved numerically for a variety of gas injection processes. Even though the method of solution is quite general, only systems with three hydrocarbon components are considered. Computer simulations of displacement processes wherein mass transfer between phases is both considered and neglected are compared, and it is shown that neglecting mass transfer can give pessimistic displacement efficiencies. INTRODUCTION The role of the gas displacement process in the recovery of petroleum has been subjected to a series of detailed analyses; as a result, a number of predictive models have been published in the literature. However, because of major simplifying assumptions, most of these models do not completely represent the physical system. As a result, the effect of making the simplifying assumptions is unknown. Therefore, a complete representation of this process — one without major simplifying assumptions —should lead to a fuller understanding of the process, and perhaps to methods of improving it. The general method of developing a model for two-phase fluid flow in a porous medium is to solve simultaneously the continuity equation, the energy equations and the equation-of-state for each phase under the prescribed initial and boundary conditions. For an isothermal system, the energy equations reduce to the momentum equation. Darcy's law. However, since natural gas is the vapor state of the reservoir liquid, interphase mass transfer may take place with concomitant changes in both the intensive and extensive thermodynamic properties of each phase. It is this phenomenon that has often been omitted in previous mathematical models. An additional relation, then, which accounts for mass transfer between the phases, must be included with the other equations to specify a complete model. Completely formulating the equations to be solved is not a difficult task1 but obtaining their solution has been intractable up to now. Availability of large-memory, high-speed digital computers now makes an attack on this formidable problem possible. This paper presents a preliminary study of the problem. Since this investigation is intended to be exploratory, it is restricted to the linear, horizontal, isothermal, two-phase viscous flow of oil and gas in an oil reservoir. In the early development of predictive models of this process, the reservoir system was considered as a unit and various forms of the material balance equation were proposed.2-4 Pressure and saturation gradients were then added in the Buckley-Leverett model.5 The Buckley-Leverett formulation considered the fluids to be incompressible; thus, the mathematical model reduces to a steady-state system. In the 1950's, studies incorporating numerical techniques were being published.6-9 These mathematical models differed in the efficiency of finite difference techniques, the inclusion or exclusion of capillarity or the number of space dimensions considered. To solve these nonlinear, partial differential equations, each phase was considered to be homogeneous with time; therefore, mass transfer between phases was neglected. The effect of mass transfer on the gas displacement process was first reported by Attra. 10 He simulated the one-dimension flow system by a series of cells in each of which the fluids were equilibrated during a time step. In addition, the pressure throughout the system during each time step was predetermined and constant-phase velocities were calculated according to the Buckley-Leverett incompressible fluid flow model. Welge et al.11 developed a model for the displacement of oil by an enriched gas where composition is considered to be a dependent variable. Fluid

By Dale A. Augenstein, S. C. Sun

A primary concern in the solution of any solid waste disposal problem is the characterization of the material. In the past, available methodology for the determination of coal refuse characteristics has been severely limited by the inseparable combination of organic and inorganic matter of which the refuse is comprised. The recent development of an electronic, low temperature asher which removes the organic material without alteration of the inorganic constituents has opened new areas of coal and coal refuse research. This study deals with the development of methodology for the characterization of Pennsylvania anthracite refuse through utilization of the low temperature asher. Seven samples of anthracite refuse were examined for verification of the methodology developed and information concerning the characteristics of the refuse.

Jan 1, 1975

By Kenneth L. Adams

10.3-1. Introduction. Equipment replacement or maintenance is of major concern to every company today. It can be the difference between an efficient and profitable operation or an inefficient and struggling one. Mining equipment, as everyone knows, will not last forever. Some equipment becomes obsolete before it wears out while some may be doing its intended job, yet not economically. It is a major problem for management to make the decision as to whether a piece of equipment should be repaired or replaced. There are numerous approaches to this problem; using formulas, charts, and graphs. These would be very good if all conditions were stable. Since conditions and requirements are constantly changing, an accurate observation cannot be derived from a set formula. A formula does, however, serve as a reminder to include most of the items that should be considered when evaluating old and new equipment. In the field of mining there are several questions which enter into the reasoning that are difficult to put into an equation. Some of these are: 1. Will the present equipment produce the required output of the mine, or an increase if necessary? 2. Are reserves adequate to support the purchase of new equipment? 3. Can the old equipment be repaired to efficiently continue the mining operation? 4. What is the stability of future markets for the product? Depreciation rate, or the write-off life, is a factor that can directly effect decisions on replacement. If the initial life expectancy is too long and the price of equipment is not written off when discarded, then the original investment has not been recovered. If the write-off time is too short, then the equipment is written off before replacement is even considered. In this case, the depreciation accrual is usually used to finance some other phase of the operation. In view of the increasing prices, the amount of depreciation fund is generally too small for much incentive to replace an older unit

Jan 1, 1968

By Garth M. Crosby, F. McIntosh Galbraith, Bronson Stringham

THE Hercules mine is located in the northeastern section of the Coeur d'Alene district, approximately 1 1/2 miles north of the town of Burke, Idaho. Surface indications of the ore deposit were first discovered in 1886, but regular mine production was not started until 1902 and was continuous until April 1925, when the known ore had been extracted. Incomplete records show that from 1912 until operations were suspended the mine produced 2 1/2 million tons of ore containing 9.4 pct lead and 7.7 oz of silver per ton, together with an estimated 2 pct zinc, 0.3 pct copper, and 20 pct iron. This operation was the first in. a series of mining enterprises culminating in October 1947 with the consolidation of Day Mines, Inc. In the same year it was decided to unwater the levels below the collar of the Hercules shaft in the hope of finding some indication of a recurrence of ore. The unwatering operation has been described in a. previous paper.' The initial exploration, following recapture of the workings, showed sufficient promise to warrant a detailed study of the mineralogy with modern techniques. The general geo1ogy of the Coeur d'Alene district, including a detailed description of the rock types encountered, has been comprehensively treated by Ransome and Calkins' in their classic paper, and only local background description, therefore, is felt to be appropriate here. The Hercules deposit transects a portion of the trough of a broad south-trending synclinorium which has been greatly complicated by faulting. More locally, it lies within a block of ground bounded on the east; by the O'Neil Gulch fault, a steep north-south overthrust of considerable magnitude, and on the west by a monzonite stock, the outcrop of which is 1/2 mile or more wide and 5 miles long. The country rock is composed of thin to medium-bedded argillites and argillaceous quartz-ites of the Prichard and Burke formations, the oldest members of the Pre-cambrian Belt Series of sediments in the area, believed to be of Algonkian age. The contact between them is a conformable gradation. The argillite is colored gray to tannish-gray and is fine-grained, compact, and generally massive in structure. Under the microscope the unaltered argillite is seen to be composed principally of anhedral quartz and a few feldspar grains which were at one time presumably partly rounded sand grains, but as a result of recrystallization and cementation by silica, the interstices are now almost obliterated and quartz grains show crenulate boundaries. The sizes of these crystals vary from 0.5 mm down to 0.1 mm in greatest dimension. In all specimens sericite comprises 10 to 20 pct of the rock and is present abundantly between most of the grains as flakes or shreds which vary considerably in size. Sometimes they form a fine felt-like mat or aggregate, and sometimes flakes are seen which appear to be good muscovite. In some specimens, separated rhombic-shaped carbonate grains are abundant, and in some instances these have been changed to sericite. Mining operations to date have explored the Hercules vein to a maximum vertical depth of 3600 ft below its outcrop, and along a maximum strike-length of 3600 ft on certain of the lower mine levels. The main orebody is irregular in outline, extending over a variable strike-length of 400 to 1500 ft; and it is intersected by a strong transverse fault that has been traced from the surface to the bottom level. This has been named the Hercules fault, and apart from the vein itself, it is the most prominent structural feature in the mine. There is good evidence that it existed prior to the introduction of ore solutions and may have influenced ore deposition, but it was also the locus of important post-ore displacement and shows a progressive right-handed horizontal component reaching 200 ft on the deeper levels. Its vertical component is not definitely known but may be considerably greater. The fault strikes 20° N to 50° E and dips westerly at angles of 70" to 45", flattening in dip where it crosses the original orebody from east to west between 1000 and 1600 ft below the surface. At about 3000 ft in depth the Hercules fault is joined by a vertical fault of similar strike, and the major post-ore dis-placement below their junction is taken up along this vertical branch of the structure, now called the Mercury fault. Recent work has been concentrated in this vicinity. Another structural feature of special geologic interest, though of little economic importance, is the occurrence of a porphyritic dike in this area. This lies a short distance above the Hercules fault, essentially parallel to it, and is 5 to 15 ft in thickness. It appears at first glance to cut the mineralization, suggesting push-apart relationship, but small stringers of the vein minerals have been observed to penetrate the dike for a matter of inches at several points. The dike is thought to be related to the monzonite intrusion. A vertical longitudinal projection of the mine is shown in Fig. 1, which illustrates most of the features discussed above. The Hercules vein was deposited along the course of a strong, persistent shear zone that now appears as a braided network of gouge seams running through more or less crushed and shattered country rock. It strikes 70° N to 80° W and dips southerly at an average of 75". Barren parts of the structure vary in width from less than 1 ft to more than 15 ft. The width of mineralized segments may be double that. Although the evidence is not conclusive, pre-mineral, normal movement along the zone may be 1000 or 1500 ft. The horizontal component is unknown. Post-ore movement appears to have been

Jan 1, 1954

By Gerald Edmunds

THE development of "fiber" or preferred orientations, during the plastic deformation of metals, and the relation of such structures to the anisotropic characteristics of worked metals has become a subject of common knowledge among metallurgists. Very little has been published, however, relative to variations existent throughout the cross-sections of rolled strips,1 and the influence that such variations may have. This paper presents such a study on a rolled zinc alloy.2 The significance of small differences in surface structure is shown in relation to bending properties. Experience with zinc and its alloys has shown that the outer radius of the bend, which is subject to tensile deformation, is the common region of failure. In considering relations of structural characteristics to bending properties it will suffice, therefore, to consider the ease and possible extent of tensile deformation resulting from various orientations. It is necessary in this connection to assume that on the average the individual crystals of a polycrystalline material act substantially as if their movements were unobstructed by the surrounding grains. This assumption is not strictly valid, but the error involved is unlikely to affect seriously the qualitative conclusions derived from its use. The mechanism of the plastic deformation of zinc crystals has been described by Mark, Polanyi and Schmid3 and by Mathewson and Phillips,4 and is regarded as ' being accomplished by block movements or

Jan 1, 1931

By Walter E. Duncan

The ores treated at the concentrating plant of the Mahoning Mining Co. at Rosiclare, Ill., come largely from the blanket replacement deposits of the northeastern part of Hardin County, Illinois, and consist of complex mixtures of galena, sphalerite and fluorspar in a gangue of quartz, limestone and sandstone. The several minerals are so intimately inter-grown that they are not liberated at the sizes usually treated by gravity concentration in the usual fluorspar plant. For this reason, an all-flotation method of concentration was developed in cooperation with the station of the U. S. Bureau of Mines at Rolla, Mo.1 This flotation separation made possible the commercial production of four coyentrates: (I) lead, (2) zinc, (3) acid-grade fluorspar and (4) metallurgical-grade Auorspar.2-3 Although fluorspar had been floated in Canada in 19214 and a plant at Rosiclare had been floating fluorspar since March 1929, fluorspar flotation was given considerable impetus by the Mahoning developments, since some features of the separation were quickly adopted by other fluorspar companies. The Bureau of Alines5 reports that in the 10 years preceding 1938 only 50,552 tons of fluorspar flotation concentrates were produced in the United States. The rise from this almost nominal figure to 153,750 tons (largely acid grade) in one year (1944)6 is indicative of the activity in fluorspar flotation occasioned by the war. The industry thus proved to be in a position to expand and keep up reasonably well with the requirements imposed upon it by such developments as occurred in high octane aviation gasoline alkylation, in the use of Freon as a refrigerant and as carrier in the Aerosol "bug bomb," and in the atomic bomb project as well as in the expansion of other activities requiring hydrofluoric acid; for example, the aluminum industry. Despite the fact that the greater portion of acid-grade fluorspar produced in the last few years went directly into war uses, the future of the acid-grade field still appears bright. As a result of the activity in fluorine chemistry during the war, it seems likely that in the relatively near future almost every phase of human endeavor will find some fluorine compound involved. The other three concentrates produced deserve some attention in these introductory remarks. The lead concentrates first produced contained both zinc and fluorspar but these did not interfere with their marketing. As will be pointed out later, the quality has steadily improved until today it is a relatively high-grade product. Although some zinc smelters accept concentrates containing limited quantities of fluorspar, it is a very objectionable impurity. Since flotation methods fail in the complete removal of the fluorspar, chemical methods were developed for

Jan 1, 1948

By Walter E. Duncan

The ores treated at the concentrating plant of the Mahoning Mining Co. at Rosiclare, Ill., come largely from the blanket replacement deposits of the northeastern part of Hardin County, Illinois, and consist of complex mixtures of galena, sphalerite and fluorspar in a gangue of quartz, limestone and sandstone. The several minerals are so intimately inter-grown that they are not liberated at the sizes usually treated by gravity concentration in the usual fluorspar plant. For this reason, an all-flotation method of concentration was developed in cooperation with the station of the U. S. Bureau of Mines at Rolla, Mo.1 This flotation separation made possible the commercial production of four coyentrates: (I) lead, (2) zinc, (3) acid-grade fluorspar and (4) metallurgical-grade Auorspar.2-3 Although fluorspar had been floated in Canada in 19214 and a plant at Rosiclare had been floating fluorspar since March 1929, fluorspar flotation was given considerable impetus by the Mahoning developments, since some features of the separation were quickly adopted by other fluorspar companies. The Bureau of Alines5 reports that in the 10 years preceding 1938 only 50,552 tons of fluorspar flotation concentrates were produced in the United States. The rise from this almost nominal figure to 153,750 tons (largely acid grade) in one year (1944)6 is indicative of the activity in fluorspar flotation occasioned by the war. The industry thus proved to be in a position to expand and keep up reasonably well with the requirements imposed upon it by such developments as occurred in high octane aviation gasoline alkylation, in the use of Freon as a refrigerant and as carrier in the Aerosol "bug bomb," and in the atomic bomb project as well as in the expansion of other activities requiring hydrofluoric acid; for example, the aluminum industry. Despite the fact that the greater portion of acid-grade fluorspar produced in the last few years went directly into war uses, the future of the acid-grade field still appears bright. As a result of the activity in fluorine chemistry during the war, it seems likely that in the relatively near future almost every phase of human endeavor will find some fluorine compound involved. The other three concentrates produced deserve some attention in these introductory remarks. The lead concentrates first produced contained both zinc and fluorspar but these did not interfere with their marketing. As will be pointed out later, the quality has steadily improved until today it is a relatively high-grade product. Although some zinc smelters accept concentrates containing limited quantities of fluorspar, it is a very objectionable impurity. Since flotation methods fail in the complete removal of the fluorspar, chemical methods were developed for

Jan 1, 1948

By Jesse O. Betterton

Recently some experimental work was conducted by the author in connection with the direct leaching of certain zinc ores with sulfuric acid with the object of subsequently recovering the zinc by electrolysis. The ores contained relatively large quantities of soluble silica, making it necessary to evolve special methods for handling the silicate solutions resulting from a leach of the ore, and also for the final removal of the silica before electrolysis. The problem was complicated by the presence of vanadium in the ore, the commercial recovery of which was very desirable. This paper, however, deals only with the silica phase of the problem. The zinc in the ore occurs in the form of a silicate (hemimorphite) and, to a very small extent, as the phosphate (hopeite) and carbonate (calamine); the vanadium is found as descloizite and vanadinite but mainly as descloizite. The experimental work was done in a plant producing 50 lb. of zinc = daily, in which the complete cycle of operations, including the final electrodeposition of the zinc, was performed continuously for one year. Some of the conclusions given here are the result of the operations in -this plant; the others are from laboratory tests inaugurated both as a guide to the plant operations and, in certain cases especially, to determine the limits in question. The practice in the 50-lb. zinc plant was to direct leach the ore crushed to pass 40 mesh, with spent electrolyte from the cells containing approximately 3 per cent. zinc and 10 per cent. free acid. This resulted in an acid leach solution and an ore residue. The acid leach solution was ' subjected to a special treatment for subsequent removal of vanadium; then the silica and vanadium in solution were precipitated together and removed as a "silica residue" by pressing through an ordinary plate and frame press. The next step was to treat the filtrate from the silica residue with zinc dust as a final purification step before it went to

Jan 1, 1923

By Jesse O. Betterton

Recently some experimental work was conducted by the author in connection with the direct leaching of certain zinc ores with sulfuric acid with the object of subsequently recovering the zinc by electrolysis. The ores contained relatively large quantities of soluble silica, making it necessary to evolve special methods for handling the silicate solutions resulting from a leach of the ore, and also for the final removal of the silica before electrolysis. The problem was complicated by the presence of vanadium in the ore, the commercial recovery of which was very desirable. This paper, however, deals only with the silica phase of the problem. The zinc in the ore occurs in the form of a silicate (hemimorphite) and, to a very small extent, as the phosphate (hopeite) and carbonate (calamine); the vanadium is found as descloizite and vanadinite but mainly as descloizite. The experimental work was done in a plant producing 50 lb. of zinc = daily, in which the complete cycle of operations, including the final electrodeposition of the zinc, was performed continuously for one year. Some of the conclusions given here are the result of the operations in -this plant; the others are from laboratory tests inaugurated both as a guide to the plant operations and, in certain cases especially, to determine the limits in question. The practice in the 50-lb. zinc plant was to direct leach the ore crushed to pass 40 mesh, with spent electrolyte from the cells containing approximately 3 per cent. zinc and 10 per cent. free acid. This resulted in an acid leach solution and an ore residue. The acid leach solution was ' subjected to a special treatment for subsequent removal of vanadium; then the silica and vanadium in solution were precipitated together and removed as a "silica residue" by pressing through an ordinary plate and frame press. The next step was to treat the filtrate from the silica residue with zinc dust as a final purification step before it went to

Jan 1, 1923

By F. L. Wolf

THE tests here described were made to obtain information regarding costs, efficiency, etc. of baking cores in an oil-fired oven, and two electric ovens, which were installed, early in 1920, in the core room of The Ohio Brass Co. and were used almost continuously during the rest of that year. Operating costs were rather high so a complete test was undertaken early in 1921. The ovens will be referred to as Nos. 1, 2, and 3. Nos. 1 and 2 are electric ovens of different makes while No. 3 is an oil-fired oven made by the firm that made No. 2. The three are equipped with similar ventilating devices. They have full-height doors at front and rear. The cores are placed on iron racks, which are lifted by Cowan trucks and rolled into the ovens through the front doors. When baked, they are removed through the rear doors. Fig. 1 shows the general plan of the electric ovens while Fig. 2 shows the plan of the oil-fired oven. Oven No. 1 is 91 in. (231 cm.) wide, 83 in. (210 cm.) high, and 79 in. (200 cm.) deep, inside measurements. The walls are constructed in, panels about 18 in. (46 cm.) wide and 4 3/16 in. (10.6 cm.) thick, packed with blocks of insulating material and covered with sheet iron. The concrete floor on which the ovens set is 13 in. (33 cm:) thick. A recess about 5 in. (13 cm.) deep was cut in the floor to provide room for the heating units. Channels and plates supported. on I beams serve as' supports for the trucks and racks. The ventilating system consists of a Sirroco No. 1 ½ fan, connected so as to draw air from the top of the-oven or from the room, through damper C, and to deliver it either to the bottom of the oven or to the floor; the delivery is controlled by the vane A. The fan and pipes are covered with ¼ n. (6.3 mm.) asbestos cement lagging.

Jan 4, 1922

By Rodney P. Smith

The rate of motion of the boundary between a single-phase a-iron region and a two phase a-iron plus ?-iron, or a-iron plus cementite region has been measured on decarburization at several temperatures between 500" and 865°C. These data, together with data for the carbon content of a-iron at the boundary, yield the diffusivity of carbon in a-iron. The activation energy for carbon diffusion is 19.8 kcal per mole for the temperature range —50° to 150°C and increases to an average value of 23.0 kcaL per mole for the temperature range 727" to 865°C. AS part of an investigation of the diffusivity of carbon in iron using a steady-state method,' the diffusivity of carbon in a-iron was determined at the temperatures 802" and 851°C. The diffusion constant thus obtained was larger, by nearly a factor of two, than that predicted from the combined results of stanley2 and of Wert.3 It seemed desirable, therefore, to continue the investigation. Since the temperature range in which the steady-state method can be conveniently used is limited, we have chosen a decarburization method, based on the laws of diffusion in metal accompanied by a phase change.4,5 If the initial carbon content of the Fe-C alloy is such that two phases coexist at the decar-burizing temperature* and decarburization is car- ried out such that the surface is maintained at substantially zero wt pct. C, the carbon distribution after a time, t, will be as shown in Fig. 1. The boundary between the single-phase and two-phase regions moves from y = 0 to y = x in the time, t, where x is proportional4 to tl". If the initial carbon (content of the sample is Ci, and CB is

Jan 1, 1962

By Charles C. Russell

Pressures developed by the coal during the coking process have been responsible for serious trouble to many companies that operate or build by-product coke ovens. The insidious nature of this trouble is indicated by the fact that oven operators, unless they have ascertained the characteristics of the coals they use, cannot determine the existence of dangerous pressures until after the damage has been done. A rather popular misconception is that ease of pushing of the charge at the completion of the coking period indicates that no dangerous pressure has been developed, but this is far from the truth; pressures developed during the coking period are often completely dissipated by the shrinkage of the charge in the completion of coking, and so any criterion based on the ease of pushing is futile. However, methods of testing coals are available that will definitely distinguish the dangerous coals from those that are safe, and such methods have been brought to a high degree of certainty. The operator should take no risk that involves either a lack of knowledge of the behavior of coals during coking or the failure to believe that certain of the coals or coal mixtures can cause damage to the oven structure. This merely tends to minimize the problem and leads to claims that coking coals of all kinds may be carbonized with no danger of damage to the oven structure provided some particular kind of byproduct oven is used. This lack of appreciation of the true situation is in itself one of the main factors leading to damage of coke-oven structures. Coals that develop more than a certain safe pressure cannot be used in by-product coke ovens without causing damage to the side flue walls. The following paper presents a highly certain method for the measurement of the pressures developed during the carbonization of coal—a method that is the culmination of many years of work on the problem in which a number of methods for the purpose were investigated and used. While actual pressures found may appear to be small when expressed in pounds per square inch, the number of square inches in each oven wall is so large that the total pressure on the oven wall becomes an enormous load.

Jan 1, 1940

By Charles C. Russell

Pressures developed by the coal during the coking process have been responsible for serious trouble to many companies that operate or build by-product coke ovens. The insidious nature of this trouble is indicated by the fact that oven operators, unless they have ascertained the characteristics of the coals they use, cannot determine the existence of dangerous pressures until after the damage has been done. A rather popular misconception is that ease of pushing of the charge at the completion of the coking period indicates that no dangerous pressure has been developed, but this is far from the truth; pressures developed during the coking period are often completely dissipated by the shrinkage of the charge in the completion of coking, and so any criterion based on the ease of pushing is futile. However, methods of testing coals are available that will definitely distinguish the dangerous coals from those that are safe, and such methods have been brought to a high degree of certainty. The operator should take no risk that involves either a lack of knowledge of the behavior of coals during coking or the failure to believe that certain of the coals or coal mixtures can cause damage to the oven structure. This merely tends to minimize the problem and leads to claims that coking coals of all kinds may be carbonized with no danger of damage to the oven structure provided some particular kind of byproduct oven is used. This lack of appreciation of the true situation is in itself one of the main factors leading to damage of coke-oven structures. Coals that develop more than a certain safe pressure cannot be used in by-product coke ovens without causing damage to the side flue walls. The following paper presents a highly certain method for the measurement of the pressures developed during the carbonization of coal—a method that is the culmination of many years of work on the problem in which a number of methods for the purpose were investigated and used. While actual pressures found may appear to be small when expressed in pounds per square inch, the number of square inches in each oven wall is so large that the total pressure on the oven wall becomes an enormous load.

Jan 1, 1940