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By R. V. Ramani, Y. J. Wang, R. Mishra

Abstract-The validity of the concept of mine resistance in ventilation systems is examined. Mine resistance is the proportionality constant relating the head generated by a fan and the quantity flowing through it when the fan is connected to a mine circuit. It is the total effect of all branches in the mine ventilation circuit. Its importance lies in the fact that for new mine projections, and in the extensions of existing mines, ventilation calculations are usually based upon adjustments of the mine

Jan 11, 1978

By Eugene Guccione

FOREWORD-In an interview published in the May 1975 issue of MINING ENGINEERING, one of the world's most respected scientists, nuclear physicist Hans Bethe discussed the reliability of nuclear power plants and their unmatched safety record. The Nobel laureate also refuted the anti-nuclear lobby's claims that nuclear plants emit harmful radioactive effluents. When asked whether the US has sufficient uranium resources, Dr. Bethe replied that "it depends on the time of introduction of the commercial and advanced breeder reactors which, in effect, can generate more nuclear fuel than they consume." The United States has enough U3O8 to last until the breeders can take over in 1995, said Dr. Bethe. But lie warned that extensive environmental damage would be caused if "environmentalists" succeed in delaying the development of the breeder. Any such delays, he noted, would vastly increase the demand for U3O8 in conventional reactors; this, in turn, would impose an immense burden on the US uranium mining industry, which would then have no alternative but to mine the low-grade Chattanooga deposits.

Jan 4, 1977

By A. J. M. Ross

BACKFILLING of stopes and other underground openings in the Homestake mine with sand tailings was undertaken primarily to reduce surface subsidence, which was wrecking much of the surface plant and a substantial part of the town, and to make possible maintenance of openings into the top of pillar stopes as mining progressed. The progress of this type of filling has been rapid and the results very satisfactory. Sand is now used in filling timbered stopes, shrinkage stopes, and abandoned drifts and crosscuts. A permanent pipe-line system for transporting sand tailings into the mine by gravity has been installed. The close filling of all stopes has resulted in important benefits other than the two major ones. The ore bodies in the Homestake mine occur as replacements in one distinctive layer of the metamorphosed pre-Cambrian sediments along a group of minor folds that together form a complex anticline that pitches at 35º to 40º to the southeast. The ore bodies are extraordinary for their magnitude. Widths of over 300 ft. have been mined and continuous ore has been followed for over 1200 ft. along the strike on some of the upper levels. Mining has been by shrinkage stoping with square-set mining of crowns and pillars. In the standard plan, shrinkage stopes 60 ft. in the direction of strike and the full width of the ore have been taken up to within 25 ft. of the level above. After filling, the crown has been mined by square-set methods, as have also the 40-ft. pillars between shrinkage stopes. In former days, when the supply of ore seemed inexhaustible, the mining of pillars was allowed to lag far behind because of the high cost. Many pillars were rather badly fractured on the level and in some places between levels. Coarse waste was used for backfilling all stopes, but in pillars it became increasingly difficult to maintain haulageways for the transportation of this fill over the top of the stopes as mining progressed The cost of maintaining these crosscuts was excessive and in many pillars practically impossible. Therefore it became the practice to fill pillar stopes by tapping the back and sides of the stopes.

Jan 1, 1939

By Charles M. Brinckerhoff

The Potrerillos mine of the Andes Copper Mining Co., a subsidiary of Anaconda Copper Mining Co., is 100 miles east of the coast town of Chanaral, in the Province of Atacama, Chile. In December 1926, the plant was placed in operation and has operated continuously ever since. Block caving is the only mining method used, as it is especially suited to the grade of ore, the position and shape of the ore bodies, and the character of the ground. At the present time there are three haulage levels, which handle oxide and sulphide ores, and production from them averages 31,000 short tons daily. The purpose of this paper is to describe the layout, equipment and systems used on the gathering or main haulage levels. Haulage Layout Because of the shapes of the ore bodies, their distances from the dumping pockets

Jan 1, 1946

By Lachlan Gilchrist

MEASUREMENTS of resistivity by methods involving the use of two current electrodes and two potential electrodes in general collinear with the current electrodes have been made extensively in geophysical investigations. In the method developed by Wenner and made use of by Gish and Rooney, and later in a modified form by Lee, the distance between the current electrodes is divided into three equal parts by the potential electrodes. In one of the methods developed by Schlumberger prospecting methods and by Koenigsberger, one current electrode is placed at considerable distance from the other and potential measurements are made in the immediate neighborhood of one of the current electrodes. The method is described as the single current electrode method, and the effect of the distant current electrode is considered negligible or a correction is made for the effect in the calculation of the resistivity. Usually the correction that is introduced is based on the calculated effect for measurements made in a homogeneous conducting material. The useful application of these methods in prospecting for minerals, of course, is in regions where nonhomogeneous conditions exist. In many of these conditions it is very doubtful whether the effect of the distant current electrode may be neglected, except for measurements in the immediate neighborhood of the "single" current electrode, and corrections based on homogeneity of the medium will necessarily be inadequate. In order to obtain a degree of symmetry that would preclude the necessity for corrections for the distant electrode and that would simplify interpretation, it appeared desirable to use a central current electrode and a set of distant end current electrodes symmetrically placed about the central electrode as shown in Fig. 1 and to provide by means of auxiliary resistances that the currents passing into the earth at the distant end electrodes should be equal. Circumstances did not permit the operation of this plan in complete detail but a degree of symmetry was obtained by using a central current electrode and two distant collinear end current electrodes and by taking measurements along a line passing through

Jan 1, 1931

By G. A. Chadwick, D. Jaffrey

The morpho1ogies and orientation relationships of the phases in the Sn-Zn and A1-A13Ni eutectic systems were examined by electron microscopy and X-ray diffraction techniques. In each alloy the "transition" from the lamellar to the fibrous morphology was found to be one of scale, not of type. The minor phase in both systems exhibited certain well developed facets which were not affected by changes in the ingot solidification rate. The crystallographic relationships displayed by the pairs of phases in both systems were also insensitive to the growth rate. In the Sn-Zn alloy, the unique relationship of: growth direction - II [1201 Sn - II [01101 Zn and ribbon interface plane 11 (101) Sn 11 (7012) Zn was determined. The Al-Al3Ni alloy phases did not possess any particular orientation relationship, though the Al3Ni phase invariably grew in the [010] direction and exhibited the same set of facet planes. These results are discussed in relation to current eutectic growth theories and explanations of the "lamellar to fibrous transition". THE lamellar to fibrous transition that occurs in many eutectic alloys has been the subject of considerable speculation and experimental study. In some alloys it can be induced solely by an increase in the solidification rate,'-3 whereas in others the transition supposedly occurs only if the lamellae are forced to grow out of the overall ingot growth direction.4-6 he cause of this latter type of transition has been attributed to deviations of the lamellae from their low energy habit planes;4'5'7 fibers are produced because the sustaining force for lamellar growth (a low energy boundary) is destroyed. Implicit in these explanations is the assumption that fibers are circular in cross-section and completely lacking in low energy inter-phase interfaces. The "natural" growth rate dependent transition has been studied less thoroughly although Tiller8 has attempted a theoretical explanation of it. Tiller's arguments are not completely satisfactory9 but his suggestion that the relative undercoolings of the solid/liquid interface for lamellar and fibrous morphologies are growth rate dependent cannot be totally discounted; it is possible, for instance, that the relative interfacial undercoolings could alter and produce the observed morphology change if the orientation relationships between the phases and the associated interphase bound- ary energies were sensitive to growth rate. Salkind et al." have reported finding a change in the orientation relationships in the A1-A13Ni system accompanying the lamellar to fibrous transition, but contradictory evidence has also been reported for this3'" and another system,12 so the position remains unclear. In an attempt to clarify matters a study was made of the "lamellar to fibrous" transition in the Sn-Zn and A1-A13Ni eutectic systems; the morphologies of these two selected systems are quite similar when viewed by optical microscopy. In the present research the morphologies and morphology changes were investigated by electron microscopy. The orientation relationships existing between the eutectic phases were also determined for both morphologies in both eutectic systems. EXPERIMENTAL PROCEDURE The materials and method of alloy preparation and ingot solidification for the Sn-Zn system have been reported previously.2 In this investigation nine horizontally grown ingots of the highest purity (99.9997 pct) were used. The temperature gradient in the melt was not intentionally varied and was approximately 10°C per cm. Seven growth rates between 1.3 cm per hr and 20 cm per hr were imposed. The A1-A13Ni alloys were prepared from "Spec. Pure" nickel and 99.995 pct aluminum by melting the components in an open alumina crucible and casting the melt into the cold graphite mold. Six ingots of the Al-Al3Ni alloy were unidirectionally solidified at growth rates from 1 cm per hr to 12 cm per hr under high purity argon. A typical ingot was 20 cm long, 1 cm wide, and 0.75 cm to 1.5 cm thick. Samples taken from the bars at positions 12 cm from the nucleation end were examined by conventional orthogonal-section metallo-graphic techniques. When required, samples were mounted for X-ray diffraction analysis using the Laue back-reflection technique with a finely focussed X-ray source. The surfaces of the A1-A13Ni specimens were prepared by mechanically polishing them down to the 1 µ diamond pad stage followed by an electropolish in 80/20 methanol/perchloric acid solution at 0°C and 20 to 30 v. The Sn-Zn specimens were repeatedly polished on an alumina pad and etched in hot dilute (2 pct) nitric acid until the diffraction spots were found to be sharp. Thin films of the alloys were prepared for observation in an electron microscope by spark machining thin discs (0.03 to 0.04 in. thick) from longitudinal and lateral sections of the bars and elec-trolytically thinning them via a jet polishing technique. For the A1-A13Ni eutectic alloy, an 80/20 mixture of ethanol/perchloric acid at 40 v and 20°C was found to be satisfactory. A solution of 70/20/10 methanol/perchloric acid/butylcellosolve at 25 v and 20°C was used on the Sn-Zn alloy. For the former alloy the jet nozzles (cathodes) and the disc clamps were of aluminum;

Jan 1, 1970

By A. V. Pegler, R. C. E. Bray

The purpose of this study is to determine the distribution of stresses in a particular portion of the ore body of an operating mine. DESCRIPTION OF GEOLOGY The ore body of this base metal mine is a lenticular mass of copper-zinc-iron sulfides, enclosed in a drag-folded quartz mica schist. The schistose host rock has steep dip which varies from north to south and an east-west strike. The drag-fold plunges eastward at about 35º, and this gives a rake of similar inclination to the bottom horizon of the ore zone. The core of massive sulfides is surrounded by an irregular envelope of disseminated sulfides in the schist, so that stope walls are comprised, in most cases, of strongly foliated schist. Adjoining the schist on the south is competent gneiss, while instrusive into the schist in a direction nearly parallel to the foliation is a series of discontinuous quartz diorite dikes, whose contacts with the schist often prove to be lines of weakness. Coarse to fine-grained pegmatites of irregular dimensions are a second type of intrusive associated with the ore body. These dikes generally strike north-east and dip westerly at about 65º. They therefore intersect the schist at an obtuse angle. The dikes are considered to be older than the massive

Jan 1, 1972

By John J. Borkert, Carl E. Gerity

In situ mining offers a new and proven method of exploiting sedimentary uranium deposits. The feasibility and economic evaluation of exploiting a uranium reserve by either open pit or underground mining should also include a similar study for in situ mining. This report compares the economics of mining several hypothetical uranium deposits at various depths, grades and reserve sizes by open pit and in situ mining. The evaluation process evolved establishing, where possible, common geological and logistical parameters to both methods of mining. The various assumptions that are unique to either open pit or in situ mining were also drafted, where possible, as equal and common functions to be used in each economic evaluation of the reserve models. After the various parameters were defined, they were programmed to an economic analysis of the various hypothetical uranium deposits by the selected mining method. From the vast array of cost evaluations, total capital and operating cost per pound U308 produced, DCF-ROI, and total net profits before State and Federal taxes were selected as the most meaningful results to compare. From the various geologic parameters used, it was determined that the economic viability of a mining venture is most sensitive to a depth-grade thickness ratio. The three selected evaluation products were plotted against the various depth-grade thickness ratios of the hypothetical uranium deposits. The results of the economic cost analyses show in situ mining generally to be a better method to maximize total net profits and to maximize return on investment.

Jan 1, 1979

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

It is an accepted procedure to represent oil field displacement problems in which the pressure is relatively constant by a potentiometric model. Variations in reservoir permeability thickness product are represented in the model by proportional variations in electrolyte depth. It is shown that significant errors are present wherever these variations are of such magnitude that the consequent differences in the electrolyte depth in the model or any regions therein are appreciable in comparison with their respective horizontal dimensions. To demonstrate the order of magnitude of this error, a hypothetical linear oil field in which permeability varies linearly with distance was treated both analytic.ally and by potentiometric model. The results obtained from the model are compared with the exact solution for various configurations of linear geometry. Deviations in flow times up to 11 per cent were observed. It is confirmed that these discrepancies can be reduced by increasing the areal scale and/or decreasing the vertical scale of the model. It is the purpose of this work to assist in the design of potentiometric models by indicating the relationship between this error and the chosen scale factors. INTRODUCTION The fundamental geometrical criterion which determines whether the pressure distribution in an oil or gas reservoir may be simulated in a conventional potentiometric model is the requirement that all reservoir parameters be constant along any vertical line between its upper and lower boundaries. This situation never obtains exactly; nevertheless, it is usually assumed that there is no permeability stratification within the section of interest, that all wells completely penetrate the sand, that there are no gravitational effects, and that variations in thickness of the sand are small compared to the horizontal extent of the field. In these ways the reservoir problem is reduced to two dimensions in which there is no variation of pressure in the vertical direction within the limits of the sand. Variations in sand thickness and/or horizontal pesmeability may be taken into account as described by Muskat, by considering the fluid conductivity of the reservoir in its horizontal plane to be proportional to the product of permeability and sand thickness. If the variations in sand thickness are small compared to the horizontal dimensions of the reservoir, then it is a good approximation to consider the reservoir problem to be essentially two dimensional. The same approximation, however, is made in the simulation of the assumed two-dimensional reservoir by the electrolytic model. The model, if it is to take into account variations of permeability-thickness product, must of necessity be made in three dimensions, that is, with variable depth. Most potentiometric model studies and theoretical treatments reported in the literature to date2,3,4,5,6,7,8 are those in which either an average permeability was used or a constant permeability was assumed. This is understandable since detailed permeability data are not usually available for the majority of reservoirs. Many investigators have recognized that the failure to incorporate permeability variations within a reservoir in model studies leads to serious error in predictions made

Jan 1, 1957

By AIME AIME

BBENJAMIN LEROY MILLER, of geology at Lehigh since 1907, is known the world around, for his former students are on every continent. He knows the earth is round for he has encircled it twice, once in 1927, and the other way around in 1937 to regain the day he lost on the first trip. An early trip was to Great Britain and Ireland, another to central Europe, a third to Spain for the International Geological Congress, a fourth to Scandinavia. His trip to the Southern Hemisphere is recorded in "Mineral Deposits of South America." In 1940 he was booked for Australia but because of the war went instead to Alaska. He spent the summer of 1941 in Central America. Thus he has gained a broad. world knowledge of mineral deposits, both metallic and nonmetallic.

Jan 1, 1942

By J. G. Montgomery

An increased drilling activity was apparent in the shallow sand fields of northwestern Pennsylvania during the year. This drilling, however, was confined to previously outlined productive areas and as new shallow sand pools of interest were discovered. Drilling for Oriskany gas proceeded at about the same rate as in previous years. One new area in Kinzua township, Warren County, was opened up during the year by a producing well, but later this well developed salt water and the productive capacities of the area are very indefinite at the present time. Production from the shallow sand areas remained approximately the same as in 1935. Production from the Oriskany or deep sand areas increased approximately 20 per cent. Two new markets were made available for the Oriskany gas during the closing months of the year by the completion of the Peoples Natural Gas Co. line from the Hebron field to Clarion County, Pa. and Penn York Gas Co. line from the Harrison field to points in northwestern New York State. With the exception of the Kinzua well, the few deep wildcat wells drilled during the year failed to discover new fields in the deep sands. Bradford Oil Field The accompanying data on the Bradford oil field include only the production obtained from the portion of that field that is in the State of Pennsylvania, for it appears probable that Cattaraugus County production, which consists mainly of the New York State portion of the Bradford field, will be included in annual oil production figures for New York State. The Bradford field production for 1936 totaled 14,533,000 bbl., and as it is necessary to estimate New York State production for the Bradford field, the accompanying data includes the production from the Pennsylvania end of the Bradford field of 13,208,000 bbl., which is an increase of approximately 8 per cent over 1935 production.

Jan 1, 1937

By M. C. Van Camp, J. D. Miller

Preliminary results are reported regarding the design and development of a pilot scale air sparged hydrocyclone for cleaning fine coal 590 pm (- 28 mesh) containing 24% ash and 1.6% sulfur. The principle of separation is the flotation of hydrophobic coal particles in the centnjugal field generated by the fluid flow in the air sparged hydrocyclone as discussed in another publication. This 152-mm (6-in.) hydrocyclone has a nominal capacity of 0.9 t/h (1 tph) and experimental results suggest that separations vastly superior to a water-only cyclone are possible. In addition the separation efficiency is as good, if not better, than that achieved with conventional flotation cells. For example, typical results indicate that 75% clean coal can be recovered at 15% ash leaving a tailing product of almost 50% ash. These experimental results coupled with the high capacity of the air sparged hydrocyclone (imagine a retention time for flotation of only two seconds compared to two minutes for conventional flotation) may represent a significant breakthrough, not only in coal preparation technology, but in the flotation of fine particles in general.

Jan 1, 1983

By Clarence Carleton Semple

At a number of mines in eastern Nicaragua, 3.5- or 5-ft. Huntington mills are used for grinding gold-ore after a pre-liminary breaking in jaw-crushers. The smaller mills are made sectional to facilitate transportation to the more difficultly accessible mines. The capacity of the smaller mills is about 20 or 25 per cent. that of the larger, but in proportion to the quantity of ore ground they require more power and are subject to greater wear. The notes herein given are from records of the performance of the larger mills at one of the mines, but, in general, apply to the smaller mills also. The Nature of the Ore. The ore occurred as a series of flat veins in a' highly-decomposed rock, probably alaskite. The primary mineralizatioll comprised pyrite, pyrrhotite, marcasite, chalcopyrite, and quartz; the oxidation-products of which were chiefly hydrated iron oxide or limonite, chrysocolla, azurite, and malachite. The kaolinization of the feldspar of the country resulted in the admixture of a great quantity of clay with the oxidized ore. The quartz was sandy, imbedded in the clay; and probably the greater part represented the residue from the crushed rock in the fissures after leaching, and kaolinization of the feldspar. There was little vein-quartz in the unaltered primary sulphide ore. Only the oxidized ore was mined for the gold that it contained in amouiits that averaged $4 per ton. No specks of gold mere ever seen in the ore, and the gold obtained by panning was always exceedingly fine. The amalgamation-bullion was about 800 fine in gold and from 60 to 80 in silver. The ore delivered at the mill contained from 5 to 15 per cent. of moisture; during the rainy season the quantity was much greater. The moisture and clay made the ore so sticky that it had to be raked over steeply-inclined grizzlies; it hung in the ore-bins,

Jan 1, 1912

Jan 1, 1909

By E. S. Moore

THE Kolar gold field has been for a long time the most important gold-producing area of India. It is situated in the State of Mysore, southern India, and not far from the City of Bangalore. The productive field is about 3 miles long and in it a gold-bearing quartz vein varying in width from 1 in. to over 10 ft. is worked. The vein carries high values to great depths and many of the mines have paid large dividends. GEOLOGICAL FORMATIONS IN THE MOLAR REGION The rocks of the Molar region are pre-Cambrian in age and bear a close resemblance to some of the pre-Cambrian formations of America. The oldest rocks seen are basic lavas, now mostly altered to hornblende-schists, with which is associated some banded iron-formation; these rocks are similar to the Keewatin schists and iron-formation in the Lake Superior region. The schist is cut by the gold-bearing quartz vein. In certain parts of the area there is also a conglomerate containing pebbles of granite, jasper,, and schist. The matrix resembles a horn-blende-schist and is intruded by small granite dikes, which in some cases have been pinched off by squeezing of the rock. This has led Dr. W. H. Smeeth, Chief Inspector of Mines of Mysore, to whom the writer is greatly indebted for information regarding this area and for his hospi-tality while visiting the field, to regard this conglomerate as a breccia due to crushing of granite dikes in the schist. Dr. J. W. Evans has regarded the conglomerate as probably a squeezed glacial boulder clay. From analogy with pre-Cambrian conglomerates in America, the writer regards it as a metamorphosed rock similar to our basal Huronian conglomerate. However, the main bearing which it has upon the problem under discussion is the apparent fact that it is younger than the schist

Jan 3, 1918

By Robert Uhley, Tsvi Meidav, L. Scharon

Before National Lead built an industrial plant on its Fredricktown property, some 100 miles south of St. Louis, a 750x500-ft area on the proposed site was investigated by electrical resistivity, seismic refraction, and churn drilling. Data were accumulated to prepare an isopach map of the overburden and to determine soundness of the dolomite bedrock and the depth to bedrock at fixed points. The site was on a gently sloping hillside. Elevations above sea level ranged from 760 ft in the northwest corner to 820 ft along the eastern edge, giving a total relief of 60 ft. There were no major drainage features, but in the northern part of the plant site there was an indicated depression area which later was interpreted as reflecting subsurface conditions in the bedrock. No rock was exposed in the entire plant site area; the surface was characterized by soil and overburden. GEOLOGY Prevailing geology in the area of investigation was relatively simple—Bonneterre dolomite overlain by residuum. The red overburden was that derived from the weathered Bonneterre. Some chert was present, interpreted in the general area as a weathering of both the Bonneterre dolomite and younger sedimentary rocks such as the Potosi dolomite. During the period of actual measurements the residuum varied from moist near surface on the western edge of the proposed site to dry near surface on the eastern edge, and for this reason overburden resistivities ranged from low on the west to high on the east. Seismic velocities were low, since the material was not well packed. The bedrock encountered was dolomite, specifically the Bonneterre of Upper Cambrian age which, according to McQueen,' consists of an upper dark gray member and a lower light gray member. These have been further divided into four zones, in ascending sequence a dark gray, sandy argillaceous dolomite; a dark brown or black, fine-grained dolomite: a light gray, finely crystalline, compact dolomite; and, at the top, a dark brown, finely crystalline, compact dolomite. In the area of investigation the upper two zones were encountered in the drilling tests as constituting the bedrock. The dolomite varies widely in porosity, which is low in the dense, fine-grained layers. In this area, therefore, both electrical resistivities and seismic velocities were high. Fracture zones in the Bonneterre are nearly Geology, and may extend to some depth, causing considerable deep weathering. It was realized that over deep fracturing the apparent resistivities

Jan 1, 1960

By Alexander B. Morris

In a recent paper on the intermittent injection of gas in gas-lift operations as opposed to continuous injection, Morgan Walker presented a comparative table showing the effect on oil and gas production from the same wells under the two methods of operation.' This table carried a column showing the formational gas-oil ratios under each method, computed by deducting the volume of input gas from the volume of trap gas. During the entire spring and summer of 1928, the present writer had been engaged on some extensive test car tests of rich gas in Glenpool, seeking, primarily, for an explanation of the apparent loss of gasoline during the summer season between the field meters and the plant master meter. Tests at several of the field meters where the richest gas was obtained regularly failed to check with other tests on the same gas made farther along the line, and the difference was not made up by the quantity of drip gasoline collected between the points. In the course of these tests, the final outcome of which is immaterial to the present purpose, an attempt was made to construct a curve representing the relationship between gasoline content of the gas in gallons per thousand cubic feet and the shrinkage in the volume of gas treated as a result of removing the condensible fractions. Such a curve was made, expressed in units of gallons per thousand cubic feet on the abcissa and the ratio vapor volume to liquid volume on the ordinate. This curve is a rectangular hyperbola and shows that 3 gal. of gasoline taken from 1000 cu. ft. of raw gas as a vapor occupied much less space, per gallon, than did 1 gal. of gasoline taken from 1000 cu. ft. of another sample of gas. This is, in general, corroborated by converting to the same units the residue settlement curves prepared by the Tidal Oil Co. and by the Natural Gasoline Manufacturers Assn., though the three curves, when plotted together, occupy entirely different positions on the paper, as shown in Fig. 1. The bearing of all this on Mr. Walker's discussion of gas-lift is this: if the formational gas is to be computed by deducting the input gas from the trap gas, some correction must be made for the vapor volume

Jan 1, 1930

By S. Fisher Morris

THE New River coal-field embraces that portion of the Appalachian coal formation which lies on the waters of the New River, principally in Fayette and Raleigh counties, West Virginia, covering a strip of territory about forty miles in length along the line of the Chesapeake and Ohio Railroad from Quinnimont to Kanawha Falls, where the New River empties into the Kanawha, the railroad following the banks of the river the whole distance. (See map.) Between Kanawha Falls and the Valley of the Ohio lies the great Kanawha region, and these two coal-fields possess a greater variety of coal than can be found in any other portion of the country, enabling them to furnish the best fuels for any of the various demands of manufacture or domestic use, the Kanawha region possessing many varieties of gas, hard splint, and cannel coal, and the New River region bituminous and semi-bituminous, steam and coking coal. The New River, in its westward course towards the Kanawha, has cut its way entirely through the Seral conglomerate or No. XII of the Pennsylvania Reports, the Umbral red shales of No. XI appearing at the foot of the mountains at the eastern end of the region at Quinnimont, one of the upper ledges of the Conglomerate forming the summits; while at the western end, at Kanawha Falls, the top of the Conglomerate is nearly down to the level of the river, a portion of the " Lower Productive Measures" forming the greater portion of the mountains. These mountains rise abruptly from the banks of the river to a height of from 800 to 1200 feet, leaving only here and there narrow strips of bottom land, a few acres in extent; and as the measures consist principally of hard sandstones, shales, and conglomerates lying in a nearly horizontal position, the mountain-sides are very precipitous, with many long and high cliffs, which, with the great height of the mountains, give the country a very rugged appearance. Along the sides of these mountains fronting on the New River and its many tributaries there are exposed the outcroppings of several veins of bituminous and, semi-bituminous coal, varying in thickness from a few inches to over four feet, five of them being workable, or

Jan 1, 1880

By Ernest R. Darby

WHEN bronze is melted in open-flame furnaces a considerable amount of slag is formed during the melting operation. This slag maybe incidental to the melting practice or it may be formed intentionally by slag-forming constituents added to the furnace charge to provide a covering for the melted metal. When this slag is removed from the furnace, it contains, in addition to the free metal carried with it by the skimming operation, a considerable amount of metal as oxide either combined or mechanically mixed with it. By the usual crushing and washing methods this metallic oxide generally goes to the tailings pile, so diluted by other foundry waste that its recovery is economically impossible. If this furnace slag is not mixed with other foundry waste, practically all of the metal oxide as well as the free metal may be recovered by some form of smelting operation. The first side of the problem to be considered is the amount of slag produced. It may not be true that the majority of foundries have no record of the amount of slag made, but it is certainly true that a very great many either have no record or pay little attention to it. Usually the collection of furnace slag for one normal week's run will be sufficient to determine the advisability of its being handled in a different manner from the crushing and concentrating generally employed. If a large representative sample of the accumulation is crushed and concentrated and the recovery compared with the results of a careful assay, it will be found that a large percentage of the total metal content was lost in the washing process. The percentage of loss will be greater than the metallic oxide content, as some free metal in extremely fine particles always is carried away in grains of slag. From such a comparison, it may be determined whether or not the slag should be melted, sold or treated as before. An example of an actual test of this kind is given below. In a foundry melting 50,000 to 80,000 lb. of bronze per day, a week's accumulation of furnace slag and ladle skimmings amounted to 16,500 lb.-over three times as much as the foundry superintendent estimated before the accumulation was made. Two tons of this material was taken as representative of the lot. This was ground through a thoroughly cleaned ball-mill and concentrated over a Wilfley table. The

Jan 1, 1928

By Charles A. Kumke

The ('slot" system of mining in use at the Golden Queen mine, Mojave, Calif., does not involve any new mining methods. It is, however, a new combination and adaptation of several stoping systems in common practice. The first plans for stoping operations at the Golden Queen mine called for ordinary shrinkage stopes. The hanging wall of the vein, where exposed, is a very hard felsite with a dip varying from 50° to 70°. The vein itself is quartz, shattered and fractured enough to make it easy to drill and blast to a clean hanging wall. There is no well defined footwall. The footwall limits had to be determined by assays as the metal content faded out into the felsite. Two shrinkage stopes, each approximately 200 ft. long, were laid out and started on the 200-ft. level. One of these stopes, varying in width from 8 to 18 ft., was carried on up to the top limits of the ore, approximately 120 ft., without a break in the operation. The shrinkage system proved entirely satisfactory and low costs resulted from this operation. In the final drawing of this stope, two timbermen followed down on the ore, scaling the hanging wall clean of small fragments and placing an occasional long stull from foot to hanging wall where it was deemed necessary. Few such stulls, however, were required. The hanging wall stood up very well until the stope was entirely emptied of ore. Waste fill was then put into the stope by the glory-hole method, through a waste raise that had previously been run to surface from the middle of the stope. The second shrinkage stope ran into difficulties about 35 ft. above the sill floor. Up to this point the maximum width of the ore had been about 15 ft. The ore suddenly widened into the footwall to a distance of 40 ft. in the north end of the stope. In extending the stoping operations over this, width, it was soon found that the vein material would not arch and support itself. Drilling in this end of the stope was then entirely too hazardous, and all operations were stopped. The ore was drawn through the chutes below as rapidly as possible and as soon as the stope was entirely

Jan 1, 1937