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By B. L. Sackett

The converting of lead matte is not a general practice at lead smelters, therefore a description of the methods used and developments made during the past 20 years in converting both lead and copper matte at the Tooele plant of the International Smelting Co. in Utah may be of interest. The Tooele converting plant consists of five electrically operated stands in which are used 96 by 150-in. horizontal, barrel-type converters each of about 10 tons capacity. The main converter building is served by a 60-ton electric traveling crane and the casting shed by a 30-ton crane. Copper matte is tapped from near-by reverberatory furnaces into launders that deliver into 8-ton ladles on the converter floor. Lead matte is brought from the blast furnaces, about li mile away, in 5-ton ladles mounted on trucks, hauled by electric locomotives. As the Tooele copper plant has never been a producer of much copper, and has been Irregular in its operation, the original converting plant with its small converters has been adcquate, although decidedly out of step with large, modern plants. Tonnage figures in this paper are for comparison with each other only, as they are not at all in line with prevailing tonnages at larger plants today. The reasons for the changes made in converting practice at Tooele, as well as the improvements in the results obtained as these changes were perfected, are given in the following pages. Lead-matte Converting, 1914 to 1927 In 1914, Oscar M. Kuchs wrote a paper for the Transactions,' describing in detail the lead-matte converting methods at Tooele up to that time. The method perfected only a few months before Mr. Kuchs' paper was written consisted, briefly, of blowing the lead matte alone, without the addition of siliceous fluxing ores, to a point where the lead and zinc were largely eliminated by volatilization. During this period

Jan 1, 1934

By Edgar Rickard

IT SHOULD be clearly understood that my remarks on economic conditions in Russia are entirely personal, and not official as an officer of the Ameri-can Relief Administration. The American Relief Administration is concerned exclusively in relief operations. The American Relief Administration has played a large part in the conditions as they exist in Russia today and it does not seem to be out of order to give you the high lights of the work accomplished. SIXTY MILLION DOLLARS FOR RUSSIAN RELIEF The Riga Agreement, by which we were enabled to carry on our operations, was signed in Riga on Aug. 20 of last year. On Sept. 8, we opened our first relief station in Petrograd. On Sept. 18, our first food was handed out in Samara on the Volga River, in the center of the so-called famine district. Altogether over 775,000 tons of food have been sent from the Atlantic Coast into Russia and we have moved this entire quantity of food over a dilapidated and wornout system of railways; and we estimate that our loss will be less than one-half of one per cent. in shortages and thefts. We have had available for this Russian work approximately $60,000,000. The bulk of this money was derived from a Congressional grant of $20,000,000 for food stuffs, a further Congressional grant of Army medical surplus supplies of $4,000,000, a grant by the Soviet Government of $11,400,000. This Soviet grant is a rather interesting phase of our operations. In all of Mr. Hoover's relief work abroad he has demanded of the country receiving relief that the people them-selves do something concrete for their own good before he would undertake to collect large sums of money in America. Secondly, before he was willing to go before Congress and ask for funds from our government he desired that the Soviet Government show their good will.

Jan 11, 1922

By J. J. Sense

10.2-1. Equipment Scheduling. In any earth-moving operation, more profit may be earned or lost by equipment scheduling than by any other single facet of the project. Many now-defunct earth movers would have survived the cost-profit squeeze had they followed an efficient, cost saving, equipment scheduling program. The ideal program would have every item of the equipment fleet working all of the time. In practice for many obvious reasons, principally mechanical conditions, operating conditions, and human nature-this ideal situation is not possible. The cost-conscious organization strives to keep as many pieces of equipment of the fleet working efficiently as much of the work day as possible. Operating efficiency requires the minimum equipment to move the predetermined tonnages in the prescribed period of time. In the case of a shovel-truck operation this would mean the scheduling of the minimum shovel and truck shifts to excavate and transport the necessary tonnage to the processing plant or waste dump. To accomplish this goal, the operator must work from the predetermined production established by the requirements of the plant, stripping ratio, mine plan, etc. This production schedule may be in terms of tons or yards per year, per month, or per hour, but must eventually be translated into tons or yards per ma- chine per unit of time. In this regard, the reader is referred to Sections 8.0, Excavation and Loading, and 9.0, Haulage and Transportation. The trend to larger and faster, and as a result more costly, equipment requires that the operator consider every possibility to keep the equipment operating around the clock, seven days per week. In some areas seven day operation may not be feasible economically, because of premium pay for Sunday work. Mine labor agreements are generally written without premium pay penalties for weekend work, and such scheduling is possible and most often desirable.

Jan 1, 1968

By V. (“Viv”) R. Forbes

[ ] INTRODUCTION "The Liberal Party believes that the State's mineral resources belong to the people of Queensland and therefore it is essential that the state, on their behalf, obtains maximum benefit from the development of those resources." (Liberal Party, 1980 "Mineral Resources Policy" p.2). "Unless sound policies are implemented.... increased production of minerals will benefit only the already wealthy and the lucky few... A resources tax is essential to ensure that above normal profits are shared by the whole community.... Labor policies on taxation must ensure an adequate return from mining developments to the Australian people." (Labor Party Policy Paper, May 1982 "National Issues relating to rapid expansion of mineral production during the eighties". p.1, 37, 47). Judging from these statements, key political parties in Australia believe that mining profits are lightly taxed and that shareholders get most of the benefits. THE REALITY The Government Share of Income statement in Table 1 attempts to measure the government share of the income generated by the "Big Five" Australian mining companies - Broken Hill Pty Ltd. (BHP), CRA Limited, Utah International, MIM Holdings Ltd., and Western Mining Corporation ("WMC"). These companies control over 70% of Australia's mining industry including all the most profitable mining operations. They dominate the mining and processing of iron and steel, aluminum, manganese, coal, oil and gas, copper, lead, zinc, silver, nickel, gold, ferro-alloys and salt in Australia. They are also major world exporters of most of these products. They are thus a representative sample of the profit-earning side of the Australian mining industry. In their two most recent financial years these five companies had income before tax of A$6,980 million. The Government share of this income was A$6,285 million or 90%. The total market capitalisation of the Big Five is about A$8,480 million. (December, 1982). Governments thus consumed the equivalent of 82% of their total value in just two years. This was taking place while the politicians drafted the above statements calling for an "adequate" share for the people.

Jan 1, 1985

By E. M. Murphy

Eight years ago the Hecla mine, a lead-silver producer, situated at Burke, Idaho, was producing ore from but two levels, the 300- and the 600-ft. All hoisting was done with a 12- by 16-in. slide-valve, double-reel, hoist. The hoist was being worked to its capacity and was unsuitable for hoisting from greater depths. As the work of opening a new level was about to be started, it became necessary to remodel the then-existing equipment or install a new hoist. The recent advent of electricity from Spokane had made available a cheaper power than steam. By a series of tests conducted throughout the Cæaur d'A1ene district, it was found that it cost $109 per year to generate a horse-power with steam. Electric power was to be had for $50 a horse-power per year. It was decided to substitute for the engines a motor-drive, of sufficient capacity to operate from the 900-ft. level, and ultimately install an entirely new hoist. After disconnecting the engines, installing a new shaft, and making a few minor changes, a 300-h-p., 2,300-volt, 514-rev. per min., variable-speed motor of the induction type was geared to the reel-shaft. The reels were supplied with 3/8- by 4-in. flat rope, and the duty of the hoist was about as follows: Cage, 1,700; ore, 2,600; car, 1,400; rope, 2,250; total, 7,950 pounds. The Bristol ammeter chart, Fig. 1, shows the combined current-consumption of the hoist- and mine-motors. The large intermittent load is clearly shown; this produced harmful fluctuations in the voltage of the long, lightly-loaded, 60,000-volt transmission-line from Spokane to the Cæmur d'A1enes. The power company received no compensation for the increased cost of producing power for this intermittent or fluctuating load. In order to protect its system from similar excessive

Jan 1, 1913

By T. Kilner, A. Plumtree

DURING an investigation of electro-deposited silver structures, samples were prepared for metallographic examination by cold mounting in plastic, grinding on successive grades of emery paper down to 3/0, and chemically polishing by swabbing using the chromic acid solution described by Gilpin and Worzala.1 Attempts to etch these samples using the standard solutions described for silver (e.g., 5 pet KCN-5 pet Na2S2O8, H2O2-NH4OH, and various potassium dichro-mate solutions) all failed to give good results due to staining of the samples and to rapid etching. The staining was caused by the chromic acid polishing solution being trapped in small fissures between the sample and the mounting plastic, and also in small fissures in the sample itself. To overcome these difficulties, an etchant was developed which prevented staining at the same time as it slowly etched the sample. This resulted in a very clean, crisply etched surface as shown in Figs. 1 and 2. This etch consisted of 10 ml of orthophosphoric acid, 10 ml of distilled water, and 2 drops of the polishing solution noted above,' which is made by mixing 100 ml of saturated CrO3 solution, 45 ml of 10 pet HC1 solution, and 800 ml of distilled water. The etching procedure involved taking the final polished sample, either immersing or swabbing in the etchant for a period of 1 to 2 min, rinsing as quickly as possible in distilled water, alcohol, and ethyl ether in that order, and then immediately drying. This etch is easily controlled due to its slow rate of etching, and produces samples with clean, sharp surfaces. Pits due to overetching produced varying contrast effects when viewed under polarized light, and are therefore obviously useful for revealing many structural details within the sample. Another effect of overetching was to produce a height varia-

Jan 1, 1968

By Thomas Downing

WHEN examining a coal property it is customary for the engineer to take channel samples at several coal faces. In doing so the extraneous matter, or partings, which can be removed by hand, or mechanically, with the facilities at hand, or proposed, is rejected. Where there is doubt concerning a stratum it is usually rejected from the main sample but part is taken for separate analysis. The samples are forwarded to a chemist for proximate analyses and determination of British thermal units, and, sometimes, ash fusion. Consideration is given all available commercial analyses but frequently the owner presents only the favorable ones. From the results thus obtained a general evaluation is made, monies are loaned or expended, markets are determined, and realization is estimated. In many instances, such grades are not substantiated and it is time we looked deeper for a basis which will prove more practical. Institute committees are working hard on analytic and use classifications and much valuable information has already been assembled and distributed. However, no single grade can be applied to any particular bed or region. Much study has been given the Pittsburgh No. 8 seam, yet that bed is subject to material differences in a single section; i. e., some strata of a face may differ from others in structure and analysis, particularly in sulfur content. From some localities it yields a good carbonization fuel, while in others it is chemically unfit for such use. Traveling south to the Logan, W. Va., field we find the Cedar Grove (Island Creek) seam a hard, gray splint coal. In the same field, and not far distant, the same bed changes, both structurally and chemically, and consists of excellent gas coal. In Illinois the moisture content of a bed may vary greatly in different localities. We have seen mines in which part of the coal is excellent for special purpose use but is spoiled by the product from another part of the same mine.

Jan 1, 1931

By H. W. Camp

The past year has had no revolutionary change in the process of petroleum refining, although there have been improvements and developments in practically every phase of operation, due largely, perhaps, to economic pressure. A few years ago, a Mid-Continent skimming plant using 35 gravity crude obtained 28 per cent. gasoline and naphtha; 9 per cent. kerosene, and the remaining yield gave gas oil and fuel oil. Figuring manufacturing cost, but not including sales, transportation, overhead and depreciation, the realization was $1.45 per bbl. of crude. Today, a thoroughly modern and up-to-date skimming plant, having all latest equipment, shows a realization of $1.74 per bbl. on the same basis. Thus it is evident that progressive developments have taken place in the last 2 or 3 years when potential earnings have been increased about 20 per cent. The average refiner is compelled to exert his utmost ingenuity in order to operate with a profitable margin. He has two factors to consider—recovery of maximum yields of profitable oils and the reduction of operating expenses. For the past 8 years, the margin between price of gasoline and cost of crude has grown narrower, constantly. Both these prices are determined by external conditions and not by the refiner himself, but the refiner might have been able to increase that margin had he more carefully controlled his output, limiting his production of light oils to current consumption. Be that as it may, this same economic pressure has given impetus to progress in improvements of primary distillation equipment, including stills, towers and heat-saving devices. A pipe still and bubble cap fractionating equipment can now replace a battery of shell stills, giving better yields, better control, elimination of rerunning and lowering of operating costs. The most recent pipe stills are being equipped either with radiant-heat furnaces or air-preheating systems and improved heat exchange equipment. These can be operated on a fuel consumption as low as 1 per cent. of the charge as compared to a fuel consumption of

Jan 1, 1929

By E. P. Mathewson

The experiments described in this paper were made at the Washoe Reduction Works, Anaconda, Mont., for the purpose of determining the relative speed of elimination of the iron, sulphur and arsenic during the process of bessemerizing copper-mattes. The samples were taken from a horizontal a barrel" converter, 8 ft. in diameter and 12 ft. 6 in. long, having an average charge of 9 tons of matte, and blown with a 16-lb. air-pressure. All samples were taken with the blast on, turning the converter from its normal blowing-position to nearly vertical, so that a conical mold, with a long handle attached to a horizontal rod, could be introduced from a platform above the converter into the mouth of the converter and a sample taken. The difficulties of taking a sample with the blast on are apparent, and the irregularities of' the lines in Figs. 1, 2 and 3 can be readily accounted for by the amount of slag in the samples. On the whole, the plotting of the results gives comparatively uniform lines, considering the conditions. The samples were taken by the regular sampling-force of the works at 10-min. intervals, as shown in Tables I., 11. and III., which also give the chemical analysis of the samples so taken. The test shown in Table I. was made with a new converter; that of Table 11. with a converter that had made three charges and had been cleaned out; that of Table 111. with a converter that had blown one charge to white metal, and had then been " washed " out and used for the test. In plotting these results, it was necessary to take, as a basis, the percentage of copper in the samples. This basis does not exactly represent the true condition, since the volume is constantly decreasing and a certain quantity of copper passes into the slag, but it is as near as can be obtained in practice. Take, for instance, in Table I., the sample No. 5, which shows: Cu, 59.9;

Jan 1, 1908

By Warren L. Howes

This paper deals primarily with an investigation of ball-mill liners that was conducted by the writer over a period of six years at the Mammoth mill in Arizona. The investigation covered a wide variety of designs of liners in both primary and secondary mills. The comparative results are presented herein as evidence of the extent to which design influenced liner life and mill performance. Flowsheet The feed to the Mammoth mill was a complex, oxidized, gold-silver-lead-molybde-num-vanadium ore with rhyolite, andesite and granite as gangue. Although the ore varied between stopes, the long-term average in character remained substantially the same. Capacity was 560 tons per day.† Fig. I shows an abbreviated flowsheet of the grinding section of the mill. Primary Ball Mill A 3-ft. Symons shorthead crusher in open circuit crushed to 1/4-in. maximum one-way dimension the feed to the 5 by Io-ft. Marcy open-end ball mill. The mill rotated at 80 per cent of critical speed and was charged with 4-in. forged steel balls. Cast balls were used for a few months at a slight reduction in ball cost, but breakage of the large balls reduced the ball size and grinding capacity to a point where the over-all costs suffered out of proportion to the saving in cost of balls. For metallurgical reasons, the finished product of the mill was maintained at minus 8-mesh. This coarse product, plus the very abrasive ore and the absence of fines in the feed, caused abnormally rapid wear of the mill liners. The resultant high operating cost led to the most extensive phase of the investigation. In order to clarify the difference in design of the various types of liners used, Fig. a shows the profiles with the arrangement in which grinding balls of maximum size would lie on the face of the liners. The results with each type of shell liner in the primary mill are shown in Table I. Since most of the testing was done with

Jan 1, 1943

By Warren L. Howes

This paper deals primarily with an investigation of ball-mill liners that was conducted by the writer over a period of six years at the Mammoth mill in Arizona. The investigation covered a wide variety of designs of liners in both primary and secondary mills. The comparative results are presented herein as evidence of the extent to which design influenced liner life and mill performance. Flowsheet The feed to the Mammoth mill was a complex, oxidized, gold-silver-lead-molybde-num-vanadium ore with rhyolite, andesite and granite as gangue. Although the ore varied between stopes, the long-term average in character remained substantially the same. Capacity was 560 tons per day.† Fig. I shows an abbreviated flowsheet of the grinding section of the mill. Primary Ball Mill A 3-ft. Symons shorthead crusher in open circuit crushed to 1/4-in. maximum one-way dimension the feed to the 5 by Io-ft. Marcy open-end ball mill. The mill rotated at 80 per cent of critical speed and was charged with 4-in. forged steel balls. Cast balls were used for a few months at a slight reduction in ball cost, but breakage of the large balls reduced the ball size and grinding capacity to a point where the over-all costs suffered out of proportion to the saving in cost of balls. For metallurgical reasons, the finished product of the mill was maintained at minus 8-mesh. This coarse product, plus the very abrasive ore and the absence of fines in the feed, caused abnormally rapid wear of the mill liners. The resultant high operating cost led to the most extensive phase of the investigation. In order to clarify the difference in design of the various types of liners used, Fig. a shows the profiles with the arrangement in which grinding balls of maximum size would lie on the face of the liners. The results with each type of shell liner in the primary mill are shown in Table I. Since most of the testing was done with

Jan 1, 1943

By James A. Barr

PUMICE and pumice are varieties of volcanic ejecta formed from lavas of the more silicic types. Chemical analyses of pumice, pumicite and perlite, from which synthetic pumice is now produced, will generally fall within the limits 65 to 75 pct silica, 9 to 20 pct alumina, less than 8 pct combined alkalies (sodium and potassium oxides), less than 3 pct calcium and magnesium oxide, less than 3 pct iron oxide, with smaller amounts of manganese, titanium, phosphorus and other oxides. Perlite is characterized chemically by having, in addition to the percentages of oxides given above, as high as 6 pct combined water. PROPERTIES Pumice is a natural silicic foam glass produced by volcanic eruptions. The molten glass foam, upon rapid cooling, trapped the steam and volcanic gases that caused it to remain extremely porous and vesicular in structure. Pumice is found in the form of large lumps several cubic feet in volume and also in grains only a few millimeters in diameter. Chemically, pumicite is of the same range of composition as pumice. The distinction between pumice and pumicite is one of particle size; the coarser materials are pumice; the fine particles, an appreciable percentage of which is in the range of 100 to 325-mesh or finer, are pumicite. If the erupting volcano ejects a silicic lava containing much gas and water vapor, the sudden release of pressure tends to free the gas. The viscosity of the lava, its rate of cooling and the amount of gas evolved, all have part in determining the degree of expansion of the glassy particles and the fineness of the vesicle. If the lava is of the correct viscosity and there is not too much gas, a pumice may be formed. Large amounts of gas may cause complete disruption into pumicite.30 In 1916, Barbour gave the name "pumicite" to the volcanic ash of Nebraska and since then the name has been widely used in industrial circles for the finer-grained commercial varieties. The glass of pumice and pumicite is transparent and colorless. It

Jan 1, 1949

By Alfred H. Brooks

PetRoleUm seepages are known in Alaska at four localities, all on Pacific seaboard. These, named from east to west, are Yakataga, Katalla on Controller Bay, Iniskin Bay on Cook Inlet, and Cold Bay on the Alaska Peninsula. Besides these, a petroleum residue has been found near Smith Bay on the north Arctic coast of the Territory. At Katalla, Cold Bay, and Iniskin Bay there has been some drilling for oil, and in the first field several productive wells have been opened up. Alaskan petroleum, so far as its composition is known, is a refining oil, with paraffin base and low sulphur content. Most of what is known of the geology of Alaska oil fields is based on the investigations of Dr. George C. Martin, of the United States Geological Survey. A, G. Maddren, of the same service, has recently made a reconnaissance of the Yakataga oil field. The data to be here presented are chiefly taken from the following publications: G. C. Martin: The Petroleum Fields of the Pacific Coast of Alaska, with an account of the Bering River coal deposits, Bulletin No. 250, U. S. Geological Survey, pp. 9 to 27 (1905). G. C. Martin: Geology and Mineral Resources of the Controller Bay Region, Alaska, Bulletin No. 335, U. S. Geological Survey, pp. 112 to 130 (1908). G. C. Martin and F. J. Katz: A Geologic Reconnaissance of the Iliamna Region, Alaska, Bulletin No. 485, U. S. Geological Survey, pp. 126 to 130 (1912). A. G. Maddren: Mineral Deposits of the Yakataga District, Bulletin No. 592, U. S. Geological Survey, pp. 143 to 147 (1914). Katalla Field The Katalla field is marked by a series of seepages and gas springs occupying an east and west belt, about 25 miles in length and from 4 to 8 miles wide. (See Fig. 1.) This zone skirts the north shore of Con-

Jan 1, 1915

By A. R. Denison

To those who are unfamiliar with the nomenclature in common use among operators in the Mid-Continent, the Permian Basin is a term to cover several widely separated productive areas. It includes what is designated in trade journals as the Panhandle fields, the West Texas fields and Southeastern New Mexico fields. It is bounded roughly on the east by the 100th parallel, on the west by the 104th parallel, 011 thr south by the Rio Grande River and various mountain ranges and on the north by the north line of the Texas Panhandle. The name is derived from the location of these three widely scattered areas which are all in a broad geosyncline which is chiefly filled with sediments of Perrliian age. The oil fields as now known and the approximate limits of the Permian Basin are shown in Fig. 1. The major part of the production in these areas comes from liniestonc or dolomite of Permian age. Some granite wash and sandstone production has been developed but ordinarily does not yield as large initial production as the limestone or dolorliite. Introduction During 1926, Texas, exclusive of the Gulf Coast, produced 12G1SS-1,-952 bbl. of oil,' of which 40,727,877 bbl., or 32 per cent., was produccd in the West Texas Permian Basin. During 1927 the same area in Texas produced 167,176,000 bbl. %f oil, of which 89,779,000 bbL13 or 52 per cent., was produced in the West Texas Permian Basin. The increase in the total production of Texas is about equivalent to the increase in production in the Permian Basin. Since production in the Panliandle District declined during the entire year, the increase of production in the West Texas Permian Basin, and also the increase in Texas, is due to the enormous increase from the West Texas fields. Daily average pl-oduction in the West Texas fields increased from 68,211 bbl. on ,Jan. 8 to

Jan 1, 1928

By John Chipman

HE lecture on 'Thermodynamic Properties of Blast Furnace Slagso prepared in 1959 and published two years later required revision in one particular before it appeared in print. The activity of SiO, was deter mined through studies of the equilibrium: SiO, (in slag) 2C (graphite) = (in Fe) 2C0 (1 atm) The ratio for each experimental temperature was found from the heat of formation of cris-tobalite, the entropies and heat capacities of the re-actants, by virtue of the third law of thermodynamics. Discrepancies between several kinds of data led to investigations regarding other equilibria involving SiO, on the suspicion that there might be an error in the third-law value of its free energy. In particular the equilibrium: was restudied by Kay and Taylor who Confirmed the earlier results of Baird and Tayloron which their SiO, activities depended. The free energy of Sic was also rechecked4 independently. These and other equilibrium measurements led to the conclusion6 that the free energy of SiO, at high temperatures is in error and requires a correction of about -5 kcal. The exact source of the error is uncertain; it could lie either in the entropy of SiO, or in its heat of formation, but it more probably lies in the latter. Adjustment of this free energy lowers the ratio by a factor of approximately 4. Accordingly it will be necessary to multiply by this factor all of the SiO, activities determined by this method.',B Referring to the lecture cited, this correction should be applied in Figs. 8, 9, 10, 19, and 20. It applies also to the experimental data of Fig. 13, bringing these as well as those on the ternary system into good agreement with the work of Taylor and associates. At the same time it eliminates the discrepancy shown in Fig. 11, lowering both lines until the upper

Jan 1, 1962

CONTENTS PAGE CARMICHAEL, NORMAN, and KIDDIE, JOHN.-Development of Mine Transportation in Clifton-Morenci District. Discussed by Norman Carmichael, Charles Legrand, B. B. Gottsberger 1 BARER, THOMAS C.-Fresnillo - Glory-hole Mining Practice. Discussed by Charles A. Mitke, B. B. Gottsberger, E. R. Thomas, Enoch Perkins..: 3 BRADLEY, P. R.-Estimation of Ore Reserves and Mining Methods in Alaska Juneau Mine. Discussed by R. M. Raymond, A. H. Rogers, Enoch Perkins, H. G. Moulton 6 BIRCH, STEPHEN.-Geology and Mining Methods of the Kennecott and Beatson Mines. Discussed by R. M. Raymond, Alan M. Bateman, Enoch Perkins 8 LEES, CHARLES E.-Mining Methods and Costs at the Iron Cap Copper Co., Copper Hill, Ariz. Discussed by A. L. Walker, A. Neustaedter 10 BELL, CHARLES N. -Mining Methods of the Telluride District. Discussed by Charles S. Hurter 10 Development of Mine Transportation in Clifton-Morenci District Discussion of the paper of NORMAN CARMICHAEL and JOHN KIDDIE, presented at the New York Meeting, February, 1924, and issued, as. Paper No. 1310-M, with MINING AND METALLURGY, March, 1924. NORMAN CARMICHAEL.-We realized the objections too the gable-bottom car, but were almost forced to its adoption by the same reasons that caused us to retain the 20-in. gage. To use another car would have meant a vast number of changes, for our transportation system was quite extensive; all mine chutes and ore bins were arranged for that car. Had we been laying out, new works, many of our decisions would have been different. MEMBER.-If you were planning it now, would you not put in box bottom cars with a rotary dump? NORMAN CARMICHAEL.-We surely would. One objection to the gable-bottom car is the spill. No, matter how carefully the cars are kept up,: there is always leakage and spill. The cost of keeping and maintaining the cars in a good state of repair, so as to keep that loss at a minimum, is not necessary for a box-bottom car.

Jan 6, 1924

By W. L. Saunders

This paper deals specifically with heading-driving as distinguished from the broader term tunnel-driving. A heading is a pilot or path-finder for the main tunnel. Some headings are complete tunnels in themselves; that is, conditions at times warrant driving a heading the full diameter of the proposed tunnel. A tunnel 10 ft. in diameter might be driven to advantage through a single heading, provided the nature of the material admits, but it usually pays to drive rock tunnels of large diameter through the heading-and-bench system. Much depends upon the material, as in soft ground tunnels of much larger diameter are driven in a single heading. In mines, the tunnels, drifts, cross-cuts, adits, etc., are usually of small diameter, hence the driving of headings applies to completed tunnels in mining-work, except, of course, in cases like drainage-tunnels and main entries leading into coal-mines, where conditions call for tunnels of large diameter, approximating those driven for railway-service. In building aqueducts the tunnels are driven from 6 to 8 ft. in diameter up to 15 or 20 ft., while in railway-service the dimensions vary from 18 to 30 ft. in diameter. There is no department of rock-excavation so difficult as that of heading-driving. This work also belongs to the most expensive class in rock-excavation. A heading is driven directly in the solid. There are no lines of lesser resistance towards which to direct the energy of the blast, but the material must be blown out by main force. The completion of the heading simplifies all the rest of the work of excavation. It is easy, for instance, to enlarge a heading either from above, below or on the sides, by breaking towards the open face through holes approximately at right angles to the axis of the heading; or the enlargement sometimes takes place by holes driven outside of the heading running approximately parallel with it, the rock

Jan 1, 1910

By Earl A. Trager

The broad picture of world petroleum production at the present time is one of an excessive supply. In any consideration of corrective measures to be applied in the future, it will be necessary to include all the producing areas of the entire world. To illustrate, United States production during 1931 was 48 million barrels1 less than 1930, whereas world production exclusive of the United States increased, owing principally to increases in Russia, Rumania and Persia. Domestic Peoduction Let us first consider the domestic production situation. Due to the fact that United States production was maintained below current demand, about 40 million barrels of oil were withdrawn from storage during 1931. Consumption increased about 1.5 per cent over 1930, which is less than the recent annual rate of increase of 8 to 11 per cent. If the machinery to prevent waste in the petroleum industry operative for more than a year will continue during the coming year and drilling is held to the 1931 level, there will be a decided improvement in the domestic situation by the close of 1932. This will be due in large measure to the drastic reduction in exploratory work and drilling which has been in effect for more than a year. Looking ahead for two or more years, probably there will be a shortage of domestic production in relation to domestic demand, unless there are discovered several large pools comparable to Oklahoma City or East Texas. Were the United States isolated, our problems in so far as petroleum production is concerned might be considered well along the road to recovery. To maintain this perspective will require the cooperation of foreign producers in the future, or protection of United Statcs markets from foreign invasion. Let us review briefly the present situation in the United States. East Texas has been the most important factor during 1931. The large initial production of individual wells and the rapid rate of drilling

Jan 1, 1932

By E. P. Jennings

These unique copper-deposits occur on the top of the Kaibab Plateau, in Cocouino county, Arizona, and extend from the northern edge of the Grand Cañon of the Colorado river to near the Utah State line. The best-developed deposits are near Jacobs Lake, 30 miles south of the Utah line, where they are 16 ft. thick, of unknown width, and are more or less continuous for 5 miles. From Jacobs Lake to the edge of the Grand Cañon, a distance of 40 miles, outcrops of ore have been found, but no explorations have been made except those of wandering cowboys and hunters. The Kaibab Plateau, the highest of the Colorado river plateaus, is a great uplift, about 90 miles long from north to south, and from 10 to 20 miles wide from east to west. Its southern face is the Grand Cañon; its western boundary a N—S. fault plane, which separates it from the Eanab Plateau, lying 1,500 ft. below. Its eastern edge is a monocline which brings the horizontal beds of the high plateau to the level of the plateaucountry of eastern Utah. The elevation of the plateau near the Colorado river is over 8,000 ft., and about 6,000 ft. at the Utah line. Its structure, as shown by the western fault, consists of about 500 ft. of Aubrey limestones underlain by 1,000 ft. of red sandstones. The limestone is the upper member of the Carboniferous, and is overlaid by the marls and shales of the Permian, 25 miles to the north on the Kanab Plateau. At the town of Kanab, near the Utah line, the Permian strata pass under the magnificent redsandstone cliffs of the Triassic, that rise to heights of 1,000 ft. above the desert. The Grand Cañon exposes a section a mile deep at the southern end of the plateau, showing all the formations from the Archean to the top of the Carboniferous.

Jan 1, 1904

By W. L. DeMoulin

A rather detailed description of the entire plant and leaching process was given in a paper recently presented to the Institute,I so this paper will cover briefly only the crushing practice of the New Cornelia Copper Co. for the year 1918. The ore, which is mined by steam shovels and loaded in side-dump cars, passes through two crushing plants. These crushing plants were constructed to crush the ore required by a leaching plant of 5000 tons daily capacity, during a crushing period of 16 hours. The primary plant reduces to less than 3 in. (76 mm.) and the secondary plant, to 1/4 in. (6.35 mm.), which is sufficiently fine for satisfactory percolation of solution through the ore in the leaching tanks, and to give-good extraction. The ore is then conveyed to the leaching tanks by a system of 28-in. (71-cm.) belt conveyors. On the way to the leaching tanks, it passes through an automatic sampling plant, where a 1-per cent. sample is taken. From the primary crushing plant, the ore is conveyed by two 36-in. (91-cm.) belt conveyors to a 10,000-ton steel storage bin, with a reinforced-concrete flat bottom. The ore discharges from the bottom of this' bin onto four 20-in. belt conveyors, which deliver it to the four units of crushers in the secondary crushing plant. There is no storage bin between the mine and the primary crushing plant. The ore breaks very coarse, is very hard, and contains a great many boulders, some as large as 4 1/2 by 4 1/2 by 10 ft. (1.3 by 1.3 by 3 m.). Jams form in the bowl of the coarse crusher in the primary crushing plant, as a consequence, and are freed by means of an immense steel hook operated from a 40-ton electric traveling crane. Primary Crushing Plant The primary plant consists of one No. 24 Gates, style K, gyratory crusher, followed by four Gates No. 8, style K, gyratory crushers. All

Jan 1, 1920