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GENERAL GEOLOGIC DESCRIPTION The Escalante silver deposit located in southwestern Utah is an epithermal vein system consisting of a hanging wall and footwall vein, straight along strike and dipping about 1.2 rad (70°). The veins consist of banded quartz and calcite fissure fillings between volcanic wall rock that are silicified for some distance outward from the vein. The walls of the veins are sharp and the silver values are irratically distributed through the various bands within the vein. Only very low grade values are to be found in the wall rock on either side. The portion of the veins above the water table have been somewhat leached, with some secondary enrichment below. The mineralogical zonation is variable due to a fluctuating water table through geologic time. SAMPLE MANAGEMENT AND COMPOSITE ASSAYS Sample data for the deposit was obtained from over 50 diamond drill holes, from channel samples cut at intervals across a drift, and from scattered surface trenches along the out- crop of the hanging wall vein. Fig. 13-1 is a surface plan map showing the vein outcrop and some diamond drill locations. Two cross sections through the vein are shown in Figs. 1 3-2 and 13-3. Nearly all of the diamond drill core of vein material was assayed in 0.3-m (I-ft) lengths, from which weighted composite assays were manually calculated for the horizontal thickness of the vein. The centers of the composited samples were plotted in north and east coordinates and by elevation, and locations and assay

Jan 1, 1980

By Mark J. Klein, A. H. Clauer

The Snoek peak induced by solute nitrogen in chromium was studied. A rapid quenching rate is required to maintain nitvogen in solution in sufficient concentrations to be detectable by internal-friction techniques. An expression for the tertninal solubility, c, of nitrogen in weight pevcent is log,,c = -4130/T + 1.05. The diffusion of nitrogen in zhromium determined by internal friction is SNOEK&apos;S theory1 of the stress-induced ordering of interstitial solute elements in the bcc lattice has been used to deduce information on the behavior of interstitials in a number of bcc metals. Under the proper conditions, internal-friction studies of "Snoek peaks" permit a determination of diffu-sivities, solubilities, and precipitation kinetics of interstitial elements. Internal-friction studies of this type have been reviewed by Nowick,&apos; Wert,= Powers and Doyle,4 and others. The behavior of interstitial atoms in Group V-A metals and iron has been rather extensively studied in this way, and results are reasonably self-consistent. However, the study of interstitial elements using internal friction in the Group VI-A metals—tungsten, molybdenum, and chromium—is much less clear. Thus far, internal-friction studies have revealed a peak in tungsten tentatively attributed to carbon5 and a peak in chromium associated with nitrogen. In chromium, Bungardt and Preisendanz6 observed a peak at 180°C for a frequency of -1 cps which they associated with nitrogen. More recently, de Morton7 determined the diffusivity of nitrogen in chromium from studies of a nitrogen peak at 160°c for a frequency of -1 cps and from elastic aftereffect measurements. The objective of this investigation was to study the nitrogen peak in greater detail, to obtain information on the behavior of nitrogen in chromium. This information is of interest because nitrogen is known to have an embrittling effect on chromium. However, efforts to deduce the role played by nitrogen in influencing the mechanical behavior of chromium are handicapped by lack of information on its behavior and distribution in chromium. EXPERIMENTAL PROCEDURES The chromium wires used in these studies were drawn from arc-melted iodide-chromium stock that had been extruded, swaged, and centerless-ground to 0.25 in.-diameter rods. Internal-friction measurements were made in an evacuated torsional pendulum similar to the one described by Kê,8 utilizing wires 0.048 to 0.030 in. in diameter. A series of analyses of the fabricated wires showed that the major impurities were approximately 100 ppm Fe and a combined interstitial content of 50 to 100 ppm C, O, N, and H. A typical interstitial analysis of the wires after processing is 15 ppm C, 2 ppm H, 50 ppm O, and < 5 ppm N. Interstitial concentrations were determined by vacuum-fusion, combustion-conductimetric, and micro-Kjeldahl analyses. In general, vacuum-fusion analysis yielded a lower nitrogen concentration than did micro-Kjeldahl analysis. For consistency, the nitrogen concentrations reported in this investigation were determined by the former method. Heat treatments were carried out by encapsulating the specimens in quartz tubes containing argon. The fabricated wires were annealed at 1150°C for 1 hr to produce uniform recrystallized grain size of about 0.015 mm. Nitrogen and carbon were added to the fabricated wires by encapsulating them in a quartz tube containing ammonia or methane, and annealing them at 1150°C for 48 hr. In some instances, carbon was also introduced by coating the wires with carbon black prior to annealing.

Jan 1, 1965

To assist the taxpayer of the oil and gas industry in correctly and expeditiously preparing his Federal tax returns, the Bureau of Internal Revenue has prepared a "Manual for the Oil and Gas Industry Under the Revenue Act of 1918." Although the endeavor has been made to anticipate all questions that may be asked, it is recognized that such a manual is only general and cannot cover all cases that may exist. The Manual is based largely upon information gathered during the fall of the year 1918 by a corps of geologists, technologists, and engineers. The investigation was undertaken primarily to furnish a basis for arriving at valuations, and depletion and depreciation deductions in connection with oil and gas properties. All fields in the United States were canvassed. Records of production of thousands of properties were collected and tabulated. These were carefully classified and studied by the most competent and experienced men in the country and the average future production curves and tables of valuation data were produced as a result of this study. Be American! Work and Save. Buy W. S. S.

Jan 7, 1919

By J. E. Schiller, S. E. Khalafalla

To improve technology for treating process water, US Bureau of Mines research has shown that magnesium oxide (MgO) has many advantages over lime or caustic soda for precipitating heavy metals. Sludge produced by MgO occupies only 0.2-0.3 times as much volume as the precipitate made wing a soluble base. While a settled, lime-formed precipitate is easily resuspended, the MgO-metal hydroxide sludge becomes cemented together on standing. Settling of the metal hydroxides from a dilute suspension is more complete than precipitates formed with other bases. Virtually any metal that can be precipitated by raising the pH can be treated using MgO. A three fold to fourfold stoichiometric excess of solid reagent is added. The mixture is reacted for five to 10 minutes. Polymer is added, and settling or filtration completes the process. Because of the greater cost of MgO compared with lime, large-scale practice of this technology will probably be limited to water containing 50 mg/L (3 gr per gal) or less of dissolved metals. For such dilute solutions, chemicals are not a large fraction of total treatment costs, so more desirable sludge properties might justify higher chemical expenses. While the MgO process is technically suitable for widespread application, the extent to which it is adopted will probably be determined by a trade-off between the greater cost of MgO compared with lime and the superior properties of the precipitates and their corresponding ultimate disposal costs.

Jan 1, 1985

By R. C. Buehl, J. P. Riott, E. P. Shoub

Pilot-plant tests have demonstrated that it is possible to produce low-sulphur sponge iron (0.03 to 0.0; per cent sulphur) as a continuous process in an internally fired rotary kiln from iron ore or mill scale; a solid carbonaceous reducing agent; and dolomite to control the sulphur content. The proportion of dolomite employed is small, so that it does not add materially to the cost of the sponge iron. The entire operation is carried out at temperatures well below the melting point of iron, so that the product is discharged in granular form. Compared with other methods for the production of sponge iron, the rotary-kiln method has the advantage of simplicity and low cost of equipment and, in addition, requires little labor for the operation of the process. The pilot-plant equipment used for this work was sufficiently large to demonstrate that the production of low-sulphur sponge iron in rotary kilns is feasible and that the costs are not excessive for applications in which sponge iron has distinct advantages over competing materials. This report constitutes a part of a program by the Bureau of Mines under a special Congressional appropriation for studying gaseous and solid fuel reduction of iron ores. Numerous articles on various phases of these investigations have appeared in Bureau of Mines publications or technical journals. Other articles are being prepared and will be published in the near future. The present program, which is under the supervision of Dr. R. S. Dean, Assistant Director of the Bureau of Mines, was initiated in 1942. However, a considerable portion of the work was a continuation of previous research performed by the Bureau. Previous investigators1-4 utilizing the rotary kiln for the production of sponge iron did not develop a suitable method for controlling the sulphur content. A review of the literature shows that the sulphur content in the product usually exceeded 0.1 per cent unless reducing agents with very low sulphur contents were employed. This high sulphur value made the product generally undesirable for use as melting stock in electric or open-hearth furnaces. The production of sponge iron in an internally fired rotary kiln has been described in detail in Bureau of Mines Bullelin 2702; consequently, this article will be confined mostly to new features of the process—that is, means for controlling the sulphur content of the sponge iron. Suitable raw materials include iron ores or mill scale; coal or coke for reducing them, and sized dolomite ( — poi-100-mesh) as the sulphur-controlling agent. A mixture of the raw materials in granular form is fed into the elevated end of the kiln, which is fired axially from the lower or discharge end. In the reducing zone, which is approximately one third the length of the kiln measured from the dis-

Jan 1, 1948

By Norman Weiss

Most papers on xanthate have dealt with principles rather than practice. On the assumption that many millmen are interested in knowing where and in what manner the xanthates are being used in mills other than their own, this paper will summarize the results of a recent survey of North American plants. A brief background story and a description of the xanthates will be followed with the results of that survey; then a few observations will be made on storage, feeding, and ore testing. The first xanthate was made in 1822 by Zeise of the University of Copenhagen. With commendable scientific detachment (or perhaps a bad cold) Zeise ignored a more pervading quality of the substance and named it for its color, after the Greek word "xanthos," meaning yellow. During the first 80 years of its recorded existence xanthate found use only in small quantity as an insecticide; after the turn of the century there was some further demand for it in the vulcanization of rubber. It was not until 1922, by coincidence the centennial of Zeise&apos;s discovery, that C. H. Keller of San Francisco, seeking sul-phidizing agents, recalled xanthate in connection with his experience years earlier when he had seen this material used in Europe to combat grape phylloxera. His successful tests with the new flotation reagent culminated in the basic xanthate patent (U. S. 1,554,216) in 1925. In the minds of many metallurgists the introduction of xanthate as a flotation reagent is identified with the beginnings of alkaline circuits, the first use of chemical collectors, and the beginning of selective flotation. Actually, these preceded xanthate by a number of years. The alkaline circuit in lead-zinc flotation was introduced by Lyster as early as 1912. In 1922 Sheridan and Griswold made their momentous discovery (Timber Butte) that alkali cyanides and zinc sulphate effectively inhibited the flotation of pyrite and sphalerite in pulps made alkaline with sodium carbonate or bicarbonate. The use of lime to retard pyrite in copper flotation began within two years after Ralston directed attention to its possibilities in 1917. These were the beginnings of modern selective flotation, which was practiced successfully on copper-iron ores at Nacozari in 1922, and at Inspiration, Cananea, and others only shortly afterward. In the lead-zinc mills, selective flotation was of even greater economic importance at that time. The Sunnyside plant of U. S. Smelting, Refining, and Mining Co. used the process earlier than 1920, and Consolidated, Timber Butte, and others soon followed. As a chemical collector xanthate was relatively a latecomer, for the Corliss patent of 1917 covered the first of these, alpha-naphthylamine, and the Perkins patent of 1919 the second, thiocarbanilid. Both of these were in extensive use before Keller&apos;s xanthate made its commercial debut in 1923. There is general agreement, nonetheless, that the introduction of xanthate was

Jan 1, 1949

By B. Amadei

INTRODUCTION Laboratory studies on the deformability and strength of intact rocks generally involve the uniaxial compression test, the standard triaxial compression test and direct or indirect tension tests. If [al ,a2 ,a3] are principal stresses [(al>a2>a3)] with compression taken as positive and tension taken as negative, the applied stress field for each one of these tests can be represented as shown in Fig. 1. However, for most in-situ conditions, such as those that exist at any point around an underground excavation, the stress field is truly three-dimensional or multiaxial [(al+ a2+a3)] and none of the tests shown in Fig. 1 can be used to predict the mechanical behavior of the rock. There are special cases for which one of the three principal stresses vanishes. Biaxial loading of rocks in-situ takes place, for instance, when rocks deform and fail in plane stress conditions. This can occur in a free unloaded surface of a rock structure such as in the wall of an underground excavation or a rock slope. Parallel to the free surface, the stress field components can be both compressive [(al>a2>0)], both tensile [(O>a2>a3)], or mixed, one component being compressive and the other one tensile [(U~>O>U~)]. The purpose of this paper is to present the results of 25 true biaxial tests that have been conducted on cubical specimens of Indiana Limestone using the multiaxial cell available at the University of Colorado at Boulder. The tests include uniaxial compression and tension, biaxial compression and combined biaxil compression- tension. Flexible fluid cushions and brush bearing platens were used for compressive and tensile loading respectively in order to minimize loading end effects on the test specimens. The paper begins with a review of previous investigations on the multiaxial behavior of rock. Then, test results are presented in terms of

Jan 1, 1984

By Oscar Kuchs

IN January and February of the current year, at the Tooele plant of the International Smelting Co., Tooele, Utah, a modification of the usual method of converting, for the treatment of copper-lead matte from lead blast- furnaces, was developed and put into successful operation. In brief, the method consists in blowing molten matte direct from the blast furnaces in basic-lined converters, without the addition of siliceous fluxing ores, to the almost complete elimination of the lead contained, and to the production of copper bullion, the resulting lead fume being recovered by the filtration of the converter gases through woolen fume bags. The process is carried out in conjunction with the converting of reverberatory copper matte, the heavy iron slag, together with the copper precipitated, being transferred to converters blowing copper matte. The purpose of this transfer is to reduce the copper oxide formed, due to the necessary overblowing of the lead-matte charge, and to utilize, as far as possible, the value of iron from the lead matte, for the fluxing of siliceous dry ores. The final slag from the copper converters is then skimmed off and transferred molten to a reverberatory smelting furnace treating copper ores. This variation from general converting practice was the outgrowth of results obtained during 1913 when it was endeavored to convert copper-lead matte in the usual manner, with the direct addition of siliceous fluxing ores. To make clear the reasons lying back of the process it is necessary to recount at some length the results of operations during that year. In the design of the lead plant it was planned to treat the copper-lead matte from the furnaces by direct converting without reconcentration and thereby to recover the copper, silver, and gold as quickly as possible. A converter plant of ample capacity, treating copper matte, already being in operation, it was only necessary: to install a bag house, a fan, and the necessary flues to complete the equipment.

Jan 7, 1914

By H. J. de Haan, L. Schenk

Scope for Thermal Recovery in Shell&apos;s Heavy Oil Fields in Venezuela The main heavy oil reservoirs on the East coast of Lake Maracaibo (Fig. I), known as "Bolivar Coast", initially contained some 20 billion bbl of oil in place with gravities in the range of 10 to 15° API. The current total production rate is about 400,000 B/D. Since the recovery to date is on an average only 12.5 percent of the initial oil in place, these fields offer a vast potential for secondary recovery methods. The reservoirs are characterized by moderate depth (generally 1,000 to 3,000 ft; maximum 5,000 ft), good formation properties such as net oil sand thickness (50 to 300 ft), high porosity (30 to 40 percent) permeability (1 to 3 darcies) and oil saturation (initially about 80 percent) and high oil viscosity (100 to 10,000 cp in situ), all of which are favorable for the application of thermal recovery processes. Since 1957, steam drive, steam "soak" and underground combustion have been tested or are being tested in this area.1-3 This paper deals with a major steam drive test in the Tia Juana field. The test was commenced in Sept., 1961, on the basis of encouraging results of laboratory experiments and pilot field tests, which will be discussed briefly. Early Laboratory Investigation of the Steam Drive Process In 1956 a series of model experiments was carried out in the Koninklijke/Shell Laboratorium, Amster- dam, to investigate the displacement of heavy oil by steam.&apos; The prototype studied was a horizontal, un-consolidated heavy oil reservoir, subjected to a linear steam drive. The model, representing a slab from this reservoir along the main flow direction, consisted of a steel tube filled with oil-saturated sand, provided with layers of rock at the top and bottom to simulate conductive heat losses to cap and base rock. Lateral heat losses, perpendicular to the main flow direction, were minimized by insulating the sides of the model. Heat flow and viscosity and gravity forces were scaled by taking the steam injection rate and sand permeability inversely proportional to the geometric reduction factor of the model, and using materials (oil, steam, cap rock, etc.) identical with those of the prototype. The main feature of the process was a frontal displacement mechanism that was observed within wide ranges of injection pressure and rate, oil viscosity, initial oil saturation and sand permeability. The high stability of this front can be attributed to the high volumetric steam flow rates, combined with the reduction in oil viscosity. High recoveries were obtained due to the low residual oil saturation of about 15 percent in the zone swept by steam. The latter value pertains to the specific test conditions (heavy oil, uncon-solidated sand) and may be lower for light oil, due to the effects of steam distillation. The applicability of this experimental approach is limited, as it does not take into account the fact that

Jan 1, 1970

By H. S. McQueen

THE central Ozark region of Missouri has inter-ested geologists and mining engineers for many years. Of particular interest are the mineralized sink-hole type deposits, some of which have produced large quantities of iron ore. Deposits of coal are also known in this part of the State, and in mode of occur-rence are similar to the iron ore deposits. Occurring in the same general area, and often associated with coal, are deposits of high-grade non-plastic flint fire clay, and high alumina or diaspore clay. At present these clays are the most important materials being mined from the sink-hole type deposits. Deposits of high-grade flint fire clay have been known to occur in this region for many years, the Kelley pit near Rolla, Phelps County, being discovered in drilling for iron ore shortly after the close of the Civil War. Mining operations had disclosed at an early date masses of rough, open-textured "sandy" clay associated with the flint fire clay. Some of the hard flint clay also con-tained rounded shot-like particles of "sand." Both types of clay were considered of inferior grade and unsuited to the manufacture of fire brick. These clays were discarded in mining as value-less, in some instances being used for road metal. Some knowledge of the high alumina content of this sandy material had been obtained by 1908 1, and by 1917 the high alumina content was gen- erally known, and commercial shipments of what is now called diaspore clay were being made. These deposits of diaspore clay associated with flint fire clay form the only known deposits of commercial importance in North America. The production of diaspore and flint fire clay at present exceeds that of any former year. Statistics covering the production and value of the former were not collected prior to 1922, and the figures given below do not cover the tonnage of clay mined by companies and used in their own plants, so that the total produc-tion probably is much larger than that indicated. The figures were collected by the U. S. Geological Survey and the U. S. Bureau of Mines in cooperation with the Missouri Bureau of Geology and Mines.

Jan 6, 1928

By A. B. Parsons

PUBLICATION of this little volume is a luxury that the Institute itself could scarcely afford. In truth, no engineer nor geologist, no assayer nor metallurgist would be likely to enhance his knowledge, improve his technique, or cut his own operating costs by reading this excellent translation of two booklets first published circa A.D. 1520. It is doubtful if the present enterprise would, in a practical or tangible way, "promote the arts and sciences connected with the economic production of the useful minerals and metals." And such, is the prescribed purpose for which the funds of the Institute should be spent. Fortunately for the Institute, the Seeley W. Mudd Memorial Fund is not so circumscribed; and the Fund Committee enjoys, greater liberty of action. The sponsorship by the Mudd Fund in 1942 of the publication of The Pirotechnia of Vannoccio Biringuccio, as translated by Dr. Smith and Mrs. Gnudi, was an innovation-indeed something of an experiment. Not only was the content of the volume significant mainly for its "cultural" and "scholarly" attributes; -but the format, an outstanding example of the fascinating art of bookmaking, bore no resemblance to the ordinary technical book. The Committee, frankly, was surprised at the enthusiastic reception accorded the volume. When a second edition was required, the Committee became convinced that the publication of material of this type, particularly if printed and bound in attractive form, was an appropriate avenue for utilizing a moderate proportion of the monies available from the Fund. The Bergbüchlein and the Probierbüchlein antedate by some years the first publication of De la Pirotechnia and of De re Metallica; but Georgius Agricola, the author of the latter work, made good use of the material in the earlier volumes. As pointed out in the Introduction, both

Jan 1, 1949

By J. P. Myers, A. H. Flagg, E. S. Mardock, J. L. P. Campbell, J. M. Terry

The development process for the use of radioactive tracers as a means of locating zones of permeability is discussed. The general techniques for the safe handling of radioactive materials is given as developed by the Lane-Wells Co. and Well Surveys, Inc. The prob1em and. successes with tracers in waterflood systetns, oil and gas injection profiles, fracture sand tracing, squeeze (ement tracing. lost circulution, and cement top localion are discussed and illustrated. INTRODUCTION The exact location of the permeable zones lying within the productive horizon has been a major problem faced by the petroleum engineer for decades, and it has been a problem which has seldom been solved with certainty. Reservoir engineers usually have to incorporate a question mark in their estimates of future reservoir performance because of blanks in the core data. Although the use of radioactivity or electrical logs for the estimation of porosity has been of great assistance in furnishing a record for the reservoir engineer, this application of logging data has not completely solved the problem. In short, the industry needs an in-situ well surveying method which can locate and estimate the permeability of the zones existing within a given pay section. The initial development work with radioactive tracers has been towards filling the need for permeability information. Progress to date permits the location of permeable zones with a qualitative estimate of their relative importance with respect to permeability. The calibration of the method is still incomplete. but present development work on tracer preparations, borehole effect correction factors, and instrument limitations give a promise of quantitative "permeability profiles" in the near future. The development work on the radioactive tracer method discussed here began several years ago in the laboratories of Well Surveys, Inc., in Tulsa. At the outset of this work, three approaches to the problem were visualized as follows: 1. A fluid carrying a soluble, chemically inert, gamma ray emitting isotope is pumped down the borehole and out into the permeable zones of the formation. The greater gamma activity of the permeable zones thus permits them to be differentiated from the impermeable zones which contain no active tracer. It was found difficult to obtain a tracer material which was chemically inert. Absorption to some degree was noted in all experiments. Also, it proved impossible to hold the tracer near the horehole where it could be logged. 2. A fluid carrying a soluble, gamma ray emitting isotope is pumped down the borehole into the permeable formations. In contrast with proposal 1, above. the tracer is chemically active with the fluids and materials in the pore spaces of the permeable zones. The permeable zones are thus indicated by their higher radioactivity level. The main difficulty in this proposal was finding a tracer material which would not be absorbed by shale as well as in the permeable porosity. Very little field work was done with this system. 3. Radioactive particles suspended within the injection fluid are pumped down the well. As the fluid enters the permeable formations, the radioactive particles are filtered out on borehole wall. The permeable zones are differentiated from the impermeable by the gamma ray activity deposited on the formation wall.

Jan 1, 1956

By John F. Burst

Nanafalia aged bauxitic kaolins which straddle the Chattahoochee River in a NE-SW trending line anchored by Eufaula, Ala., and Andersonville, Ga., apparently derive from the same Piedmont source. A genetic interpretation is offered which describes the Nanafalia as a shoreline deposit in which the blanket elastics of the Andersonville area represent an ancient freshwater swamp and the lenticular Eufaula bauxites represent meander fills in an ancient tidal flat or barrier island. Mining operations in both areas are described and economic forecasts are presented.

Jan 1, 1975

By C. Feng, C. Elbaum

AN investigation of the fracture plane of titanium, containing 0.5 wt pct to 1 wt pct 0, was recently carried out by fracturing large-grain specimens in a tensile jig at room and liquid air temperature. The experiments were carried out on strip specimens 3 to 5 mm wide, 2 mm thick and about 100 mm long. The specimens consisted of several grains, each occupying the entire cross section of the strip and extending over a length varying from 5 to 20 mm. These specimens were prepared by a method described elsewhere.&apos; The identification of the fracture plane was possible in twelve of approximately thirty crystals investigated. In some cases severe local distortion took place near the fracture and prevented identification of the fracture plane. On, at least, three occasions the fracture propagated in a helicoidal

Jan 1, 1959

By F. H. Deily, H. W. Lownie, E. A. Loria, L. W. Ledgerwood, M. W. 207-000-000-012 Mallett, J. E. Eckel

MANY investigations have been made of the residual gas content of iron and steel. Very few, if any, quantitative measurements have been made on the oxygen content of cupola-melted iron.&apos; The fact that cast iron is a complex alloy rather than

Jan 1, 1957

By L. D. Jaffe, D. C. Buffum, F. L. Carr, M. Goldman

IN an earlier investigation by Jaffe and Buffum,&apos; a time-temperature transformation diagram for isothermal temper embrittlement of an SAE 3140 steel showed some suggestion of two overlapping "noses." These authors mentioned, as a possible explanation, that their specimens embrittled at 525°C were slightly softer than expected, perhaps because of some slight unknown variation in tempering treatment. Because a double nose would be of considerable importance in elucidating the mechanism of the temper brittleness reaction,&apos; it appeared worthwhile to redetermine the portion of the diagram near 525°C. The same heat of SAE 3140 steel was used as in the previous work, and experimental procedures were the same.&apos; Blanks were austenitized 1 hr at 900°C, water-quenched, tempered 1 hr at 675°C, and water-quenched. Groups were then embrittled isothermally at the temperatures and for the times shown in Table I, and water-quenched. For comparison, one series was tempered various times at 675°C, water-quenched, and given no further treatment. V-notched Charpy impact specimens machined from the blanks were tested over a range of temperatures. Energy absorption and percentage of fibrous fracture were plotted vs test temperature for each group, utilizing both the new data and those for the same treatment obtained in the earlier work,&apos; except that the earlier data at 525°C were discarded.* are given in Table I. After a small correction (-0.28°C/BHN) for the variation of hardness" from Rockwell C24, the transition temperatures were plotted against embrittling time for each embrittling temperature, using the data of Table I for the range of treatments covered there, and using previous dataL for other treatments. By cross-plotting, Fig. 1 was finally obtained. It appears from this figure that if there are two noses to the time-temperature-transformation diagram for isothermal temper embrittlement of SAE 3140 steel, the upper nose must lie above 575°C. As interest had been primarily in the region near 525°C, a thorough study of embrittlement above 575°C was not made. Fig. 1 shows that such a study is needed, especially since preliminary measurements at the tempering temperature of 675°C reveal much more rapid embrittlement than would be expected from the data at lower temperatures. The transition temperature increased, approximately, as the logarithm of the tempering time (about 81/2°C, after correction for hardness, for each doubling of the time).& Since three tensile specimens tempered 240 hr at 675°C and water-quenched had an average yield-tensile ratio of 0.844, with 70.7 pct reduction of area, it is assumed1 that the tempering temperature is not above Ae1. The omission of a quench-and-reheat operation below Ae1, for the groups tempered at 675°C, as compared to groups embrittled at lower temperatures, should not affect results." More detailed investigation of embrittlement above 575°C is now under way.

Jan 1, 1954

By A. B. Stevens

IN the calculation of oil and gas reserves by volumetric methods, it is necessary to know the volume of the reservoir available for the storage of the oil or gas. To obtain this information, a number of methods for determining both total and effective porosity of reservoir rocks are used. For the determination of the total porosity, or the porosity that considers all of the pore spaces, connected and isolated, the Russell7,8,9 and the Melcher-Nutting1-6 methods are the best known and most- widely used in the petroleum industry. These methods have disadvantages in that they give the total porosity of the samples and in that the samples are reduced to grain size in the determination. In many sands there may be little or no difference between total and effective porosity but in other sands there may be considerable difference between the two. For the estimation of oil-and gas reserves the effective porosity of the reservoir rock always should be used. Of the methods that have been developed for the determination of the effective porosity of reservoir rocks, or the porosity that considers only those pore spaces that are connected, the best known are the so-called Boyle&apos;s law6,10-13, and the Washburn-Buntings.5,6,11 methods. The Wash-burn-Bunting method is probably the most widely used because it gives effective porosity, it is rapid, and the apparatus is inexpensive. This method employs a special porosimeter in which the core sample is placed in direct contact with mercury. This method has disadvantages in that, as a result of the high vacuum created in the pore spaces, mercury is drawn into the openings, making it impossible to use the sample for a check determination and also ruining the sample for any further test, such as permeability. The purpose of this paper is to describe an appa-ratus employing the same principle as the Washburn-Bunting method but so designed that the core sample does not come in contact with the mercury in the porosimeter.

Jan 1, 1939

By R. W. Raymond

I DESIRE to call the attention of the Institute briefly, and by no means in the way of an exhaustive description, to the interesting developments recently made on the east bank of the Hudson River, in the mining of spathic iron ores. The mines to which I refer are those of the Hudson River Spathic Iron Ore Company, situated at Oakhill on Hudson, Columbia County, N. Y., five or six miles from the city of Hudson, and opposite the town of Catskill. The ore occurs in beds, constituting apparently members of the stratified series of rocks on that side of the river, coursing nearly north and south, parallel with the river, and dipping east. The principal out- crops are presented in a range of low hills, about one mile from the river, along the crest, on the western face of which the stratum of spathic ore is more or less exposed. The length of the line in this range, along which, at various openings the ore has been discovered in profitable quantity, is about two miles. To the northward, the hills are somewhat higher, and the stratification may possibly be disturbed, though the rapid examination of the locality which I made in November last did not, include any determination of this

Jan 1, 1876

By Waldemar Dyrssen

I CANNOT agree with the results obtained by the committee. The gas obtained per pound of coal is too high in calorific value and the amount of gas is too large. It should require about 190 cu. ft. of oxygen per 1000 cu. ft. of gas and not 75.14 cu. ft. as is given on page 35 of the report, or about 2.5 as much. Calculating the calorific value per pound of coal from the coal analysis given on page 34, according to Dulong&apos;s formula, the gross and net calorific heat are about 12,890 and 12,430 B.t.u., respectively, which is about 18,600 and 17,900 B.t.u., respectively, per pound of carbon in the coal. This is an unusually high figure, as ordinary bituminous coal has less than 18,000 gross B.t.u. per pound of carbon. Assuming, however, that the calorific values calculated are correct, the following results are obtained. On page 35, the net calorific heat in the gas per pound of coal is given as 12,856 B.t.u. and calculation for the gross value shows it to be about 13,950; in other words, the gas contains (13,950 - 12,890) ÷ 12,890 = 8.2 per cent. more heat than the coal on a gross basis and (12,856 - 12,430) ÷ 12,430 = 3.4 per cent. more on a net basis. The calculated amount of carbon in this gas per pound of coal gasified is about 0.688 lb., or nearly the same as in the coal (0.6972) and the gross and net calorific heat per pound of carbon in the gas are 20,300 and 18,710 B.t.u. which are 9.1 and 4.5 per cent. more than for the carbon in the coal. These calculations show that the results given in the report cannot be correct. There must always be some heat losses in the gasification of coal, in radiation from the producer, and in sensible heat in the gas; there must be some loss of carbon in ashes and in soot, and there are always tarry vapors in the gases that are condensed and do not appear in the fixed gases. From the excessive heat in gas per pound of carbon contained therein, it can also be concluded that the percentage of hydrogen in the gas is too high, from which it follows that the moisture decomposed is too high and that more free oxygen must be used.

Jan 11, 1924

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a &apos;moral victory&apos; as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp&apos;s brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953