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By Warren B. Davis

Introduction If a subject this broad were to be covered in even moderate detail, it would require a set of books about the size of an encyclopedia. Since an acceptable length for this paper is a small fraction of one book, it will have to consist of a few samples of the subjects that I think are the most pertinent to World Fossil Fuel Economics. I plan to discuss the trends in world energy demand, particularly in the U. S. and Europe; the consumption patterns and cost patterns of oil, gas and coal; the prospects of synthetics; and some of the effects of government regulation of fuel economics. World Energy Demands Fig. 1 shows Free World energy consumption by principal areas for the period from 1950 to 1970 and a forecast for the period from 1970 to 1985. These figures include only the measurable forms of energy. An example of energy consumption that is not measured is the wood a peasant might cut in the forest and bum in a fireplace to heat his home. This omission probably is a very small part of the total. This forecast is not identical with that of Wayne E. Glenn (see Page 22), but the differences in total energy consumption are not significant. The U. S. consumption, which was 55 percent of the world total in 1950, had declined to 45 percent by 1970, and is expected to decline further to 38 percent by 1985. Western Europe has consumed 28 percent of the total from 1950 to 1970 and is expected to lose share slightly between 1970 and 1985. The most spectacular gainer is Japan, which increased consumption from 2.5 percent in 1950 to 7.5 percent in 1970, and is expected to reach nearly 17 percent by 1985. The over-all Free World total has grown from 62 quadrillion Btu (equivalent to 29 million B/D of crude oil) in 1950 to 151 quadrillion by 1970, an average growth rate of 4.5 percent/year. This total is expected to grow at a rate of 5 percent/year between now and 1985, reaching 315 quadrillion Btu (equivalent to nearly 150 million B/D). Fig. 2 shows which fuels supplied this energy between 1950 and 1970, and gives an estimate of how it will be supplied between 1970 and 1985.

Jan 1, 1971

By N. J. Grant, K. M. Zwilsky, J. T. Blucher

DISPERSION strengthening has been studied in a number of metal-metal oxide systems. To date those pure metals available as fine particles (less than about 5 p) have received most of the attention whereas solid solution alloys, generally unavailable as fine powders, have not been examined in detail. The present work was undertaken as a preliminary study to evaluate combined effects of solid solution and dispersion strengthening. The technique used was that of dry mixing of elemental powders with the oxide, followed by a diffusion treatment to obtain the solid solution. The average particle size of the powders used was: Ni, 3.2 p (1nco B); Mo, 1.3 p (Fansteel Metallurgical Corp.); and y alumina, 0.027 \i (G. L. Cabot, Inc.). The particle size of the metal powders was determined by Fisher subsieve measurements while the size of the alumina was calculated from a surface area determination. Prior diffusion studies on small compacts pressed from mixtures of nickel and molybdenum powders with and without 7.5 vol pct A1203 indicated that diffusion was complete after sintering for 100 hr at 1800" , or 50 hr at 2250", or 20 hr at 2550° F. The progress of diffusion was followed by microstructural examination and micro-hardness measurements. Two observations made during the diffusion experiments were: 1) That diffusion proceeded faster in the presence of alumina, confirming similar findings by Seith and Loepman on "SAP'';4 2) That diffusion at temperatures above 1800°F resulted in coarsening of the alumina in agreement with results of other authors reporting on the growth of alumina in nickel above 1800"F. Similar experiments in the Ni-Cr system using 6 and 20 p Cr powders of 0.7 and 0.025 pct 0, respectively, did not result in significant solid solution formation even after long time exposures at 2550° F. It must be concluded that chromium oxide films around individual particles interfere with normal diffusion. A nickel alloy containing 5 wt pct Mo and 7.5 vol pct alumina was fabricated by dry mixing of powders, isostatically compacting at 30,000 psi, diffusing for 120 hr in dry hydrogen at 1800" F, and extruding at 1950"~ with a 30 to 1 reduction of area ratio. The resulting structure, presented in Fig. 1

Jan 1, 1963

By W. H. Morris

I have been requested to present a paper on the slag from the basic Bessemer process, as prepared for fertilizing. Since Professor W. B. Phillips presented at the Birmingham meeting, in May, 1888, an able paper* on this subject, as it had been developed up to that time, there is little for me to add. Perhaps later practice may be best expressed by the homely proverb, " The proof of the pudding is in the eating." As the Pottstown Iron Company is the only concern manufacturing basic Bessemer slag in this country, the writer's personal experience in the use of it may be interesting. Our slag averages: P2O5, 21.37; SiO2,, 5.10; MgO, 5.90; Al2O3, 4.01; MnO, 5.56 ; FeO, 12.00; Na2O and K2O, 0.80; CaO, 45.26; total, 100.00. This gives over 76 per cent. of phosphoric acid and lime. In using this material for several years past on my lawn and garden, I have found it better than anything else I have ever tried ; and I am sure that any of you who had seen my green grass last month (January), wherever the ground was bare from snow, as bright as in early fall, would be forced to acknowledge the advantages of this soil-enricher. The grass on my private lawn, and around the company's office, is green the whole winter; and we have raised very good sod, by the use of this phosphate, where, without it, we were unable to get the grass to grow either from seed or by repeatedly sodding. The same statement applies to garden-truck; when the soil was first broken up, a good crop of vegetables was easily raised. My own gardener has had several premiums from Philadelphia seedsmen (notably, last year, the first premium for sugarcorn raised in the first part of July, being the first sample sent them). Our method of applying this material is to grind the slag very finely, so that at least 60 per cent. of it will go through a 150-mesh screen (i.e., say, 22,500 meshes to the square inch), and 90 per cent. through a 100-mesh screen (i.e., 10,000 meshes to the square inch). This fine grinding seems to be the main condition of success.

Jan 1, 1893

By Hans W. Schreiber, David W. Neuhaus

INTRODUCTION The Sunshine Mining Company's 16-to-1 silver deposit and mine project are located 384 kilometers (240 miles) southeast of Reno and 360 kilometers (225 miles northwest of Las Vegas, near Silver Peak in Esmeralda County, Nevada, at an elevation of about 2,300 meters (7,000 feet). The deposit was being developed, in 1980, with an anticipated mining and milling rate of 150,000 tons per year (500 tons per day). Mining is by underground, blast-hole open stoping with rubber-tire haul-age out of the mine to a straight cyanide plant about 5.5 kilometers (3.5 miles) from the mine. The product is a gold-silver dore which may be further refined at Sunshine's new refinery in Kellogg, Idaho. The cash required for investment was obtained through the sale of $25 million worth of 6 l/2h silver-indexed bonds. Some $15.0 million were directed toward development of the 16-to-1 Mine. These bonds, which were underwritten by Drexel Burnham Lambert of Chicago in April of 1980, are redeemable on or after April 15, 1985 or payable at maturity on April 15, 1995 at the greater of either $1,000 or the market price of 50 ounces of silver. Behre Dolbear acknowledges the kindness of the Sunshine Mining Company (Sunshine) for allowing the presentation of the subject material for this Case Study. APPLICABILITY AS CASE STUDY Reasons for Choosing 16-to-1 The reasons for choosing the 16-to-1 Mine as a case study follow: 1. The project is simple and compact. 2. The mine is small -- thus relevant to other contemplated precious metal operaions in the western states. 3. The areas of risk are readily identifiable. 4. A change to project financing was under consideration. 5. A substantial decline in the price of silver was experienced during development. 6. Behre Dolbear undertook one of the first feasibility studies of The 16-to-1 and thus has some familiarity with the project. Historical Summary The following history serves as a setting: 1. Late 1920s -- minor "production"(?) via shallow pitting. 2. Mid-1930s -- neighboring Nivlok Mine successfully put into production. 3. Late 1950s (?) -- Callahan Mining Company acquired rights and drilled -- results inconclusive; Callahan turned property back to Arthur Baker, who had originally interested Callahan in 16-to-1 on the basis of its similarity with Nivlok Mine.

Jan 1, 1985

By P. W. Davies, B. Wilshire

PREVIOUS investigations on the effect of stress changes on the high-temperature creep and fracture behavior of metals have been confined mainly to the testing of complex alloys.172 Most of these alloys are structurally unstable during creep because of such effects as recrystallization, overaging, and so forth, so that the results are of limited application. The only pure metals on which observations have been made appear to be lead3,4 and aluminum5 tested under conditions which allow rapid recovery to take place. The present work describes the effect of a single stress change on the creep and fracture properties of a metal in which recovery is very slow. A Ni 0.1 at. pct Pd alloy was used so that grain growth and recrystallization did not occur during creep testing. Specimens of this material were heat-treated to give a uniform grain size of 10 grains per mm. Creep tests were carried out at 500°C using Denison T.47 machines; details of the creep-testing procedure have been given elsewhere.' All specimens were tested at a constant initial stress of 5 tpsi which gives, in an uninterrupted test, a minimum creep rate of ~10-3 pct per rnin and a life of -7000 min. Specimens were loaded under these standard conditions and after 1500 min, when secondary creep was well-established, the stress was changed to either 8, 7, 6, 4, 2, or 0 tpsi for a period of 180 min. After this time the original load was reapplied and the tests continued to fracture. The results are shown in Fig. 1. No detectable creep occurred during the period of reduced load, even at 4 tpsi, and on reverting to 5 tpsi the creep rate returned to its original value immediately. Increasing the stress to 8, 7, or 6 tpsi resulted in an instantaneous extension followed by a period of transient creep leading to a new steady creep rate. On returning from these higher stresses, the resulting creep rate at 5 tpsi was very low compared with the original value. Furthermore, the higher the stress during the stress change, the lower was the final creep rate. It is significant that the rupture life is comparatively unaffected by any of these stress changes.

Jan 1, 1965

Discussion of the paper of Edward Keller, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 80, August, 1913, pp. 2093 to 2115. GEORGE L. HEATH, Hubbell, Mich.:-In the admirable paper quoted, is mentioned the date when the 5-g. assay was first proposed is a standard for refined metal. The author omitted, however, to mention a later paper,1 which contains a proposal for the use of a single " stock solution " of mixed acids, which has been adopted by the American Brass Co. and others. The suggestion is made that such a scheme would save time and manipulation with converter metal. Unless the low current of 0.5 ampere and the large excess of nitric acid, recommended by Mr. Keller, have been proved necessary to hold up antimony-and selenium, the use of the mixed solution which we have adopted will permit the use of 1 ampere of current per square decimeter of cathode surface, until the solutions have become colorless. At this point, we wash the covers and reduce to 0.5 ampere to prevent oxidation or contamination. The ordinary split electrodes have been used. With strong current, at least, the use of an excess of nitric acid tends to the neutralization of the solution before the end of the period, from the formation of too much ammonia., and this is detrimental. Antimony deposits more easily than arsenic, and tends to drag the latter dorm with it. There must, be some such explanation for the fact. that we have made hundreds of deposits of pure cathodes from Lake material containing no antimony, but as. high as 0.5 per cent. of arsenic. By placing an assay beaker in the Frary solenoid, or rotary device, a cathode deposit with less than 0.01 per cent. of impurity is easily obtained in one operation from 3 g. of whitneyite, carrying from 7 to 10 per cent. of arsenic. In such a case, it is necessary to test several times for end point and remove the plate as soon as no brown tint appears (at once) when hydrogen sulphide water is added to the test portion on the spot plate. Edward Keller's proportion,, of solvent figure out as 14 cc. of strong nitric to 11.4 cc. of sulphuric acid (1.84).

Jan 11, 1913

By K. R. Garr, D. Kramer, A. G. Pard

EXPOSURE of steels to a fast-neutron flux results in helium generation by (n,a) reactions. The elements Fe, Ni, Cr, and N are major participants in the (n,a) reaction and helium concentrations of up to 5 x 10"5 atom fraction could accrue in fuel clads after a year of service in proposed fast reactors.' Such concentrations of helium can severely reduce the ductility of alloys at temperatures of reactor service. The phenomenon of helium embrittlement has been studied in austenitic alloys which were irradiated with particles as a rapid means of producing significant helium concentrations.2-5 The pres-ent work is a continuation of those studies and is concerned with helium effects in a ferritic stainless steel. I) EXPERIMENT Small sheet tensile samples with a gage length of 0.50 in. and a cross-section of 0.009 by 0.040 in. in the gage length were punched from cold-rolled type 405 stainless steel. The composition of the steel used in this work is given in Table I. Prior to helium injection, the samples were annealed 24 hr at 760°C which caused complete recrys-tallization with a resulting grain size of 40 u. This treatment rendered the alloy completely ferritic and carbide particles, 0.5 to 2 uin size, were present at a density of 1 x 1011 cm-3, Fig. 1. These carbide particles were probably a mixture of the M7C3 and M4C types.6 Helium was injected into the samples by irradiation with a particles from a cyclotron, which produced uniform concentrations through the center two-thirds of the sample thickness. Details of this procedure were previously described.' The average helium concentration, as measured on a gas-source mass spec- 11) RESULTS A) Tensile Tests. The results of the tensile tests are given in Table I1 where the values shown are averages of two samples. There was no significant effect of helium on either the yield strength or tensile strength at 550" or 650°C but an increase in both is apparent at 750°C. Samples with helium showed a 20 pct loss in total elongation at 650°C which increased to a 55 pct loss at 750°C. The uniform elongation is not affected by helium at 550" or 650°C but shows an increase at 750°C.

Jan 1, 1970

By P. E. Fitzgerald, J. C. Gindlesperger

1. The Acidizer (Trade Publication of Dowell Inc.): No. 1. Aspects of Acidizing. No. 2. Inhibited Acid. No. 3. The Value of Acidizing. No. 4. The Chemistry of Acidizing. No. 5. Jelly Seal in Acidizing. No. 6. Loosening Stuck Drill Pipe. No. 7. The Geology of Acidizing. 2. J. C. Albright: Acid Is Being Used for Cleaning Intake Wells in Water-flooded Areas. Oil Weekly (Sept. 23, 1935) 24. 3. D. H. Bancroft: Acid Tests Increase Production of Zwolle Wells, Miller County, Arkansas. Oil and Gas Jnl. (Dec. 22, 1932) 42. 4. A. H. Bell: Possible Areas for Acid Treatment in Illinois. Pub. of Illinois- Indiana Petr. Assn. and lllinois Geol. Survey (1934) 49. 5. N. A. Best: Treating Limestone Formations Consists of More than Pouring Hydrochloric Acid in the Well. Oil and Gas Jnl. (April 5, 1934) 54. 6. L. G. E. Bignell: Sealing Off Salt Water from Oil-bearing Formations. Oil and Gas Jnl. (Jan. 30, 1936) 59. 7. L. G. E, Bignell: Acidizing Upper Limestone Formation Facilitated by the Use of Jelly Seal. Oil and Gas Jnl. (March 5, 1936) 30. 8. L. G. E. Bignell: Bailing, Shooting, Tubing and Acidizing Have All Been Tried in Deep Test at Binger, Oklahoma. Oil and Gas Jnl. (May 10, 1934) 38. 9. L. G. E. Bignell: Dowel1 Incorporated Increasing Its Field Service for Acid Treatment of Wells. Oil and Gas Jnl. (June 7, 1934) 62. 10. L. C. Chamberlain, Jr. and P. E. Fitzgerald: Oil Well Acidizing in Michigan. Mich. Acad. Sci. Papers (March 1937). 11. M. E. Chapman: Some of the Theoretical and Practical Aspects of the Acid Treatment of Limestone Wells. Oil and Gas Jnl. (Oct. 12, 1933) 10. 12. C. E. Clason and J. G. Staudt: Limestone Reservoir Rocks of Kansas React Favorably to Acid Treatment. Oil and Gas Jnl. (April 25, 1935) 53. 13. W. F. Cloud: Cleaning and Reconditioning Oil Wells—Reconditioning by Use of Acid (Petroleum Production 1937), 389. University of Oklahoma Press, 1937,' 14. K. A. Covell: Developments in Eastern Production Practice. Acid Treatment of Michigan Wells. Amer. Petr. Inst. Prod. Bull. 213 (1934) 77; Oil and Gas Jnl. (May 31, 1934) 89; Oil Weekly (May 28, 1934) 17. 15. B. A. Cunningham: Acid Treatment of Lime Formations. Amer. Petr. Inst. Paper 901 (1933) 9D, 1-7. 16. 0. I. DeSchon: Pondera Production up 22,000 Barrels. Montana Oil & Min. Jnl. (Jan. 26, 1935) 1.

Jan 1, 1937

By P. D. Merica

The remarkable phenomena exhibited by the aluminum alloy known as duralumin were discovered during the years 1903-1911 by A. Wilm1,2 and have been described by him and by others. 3 4 5 6 The unusual feature of this alloy is the fact, as was shown by Wilm, that it can be hardened quite appreciably by quenching from temperatures below its melting point followed by aging at ordinary temperatures, which consists merely f allowing the material to stand at these temperatures. The hardness is not produced by the quenching alone but increases during the period of aging, which may be from one to three days. Cohn3, 5 gives data showing the increase of hardness of duralumin during aging, after quenching in water from about 450° C. Upon annealing, the alloy so hardened by aging is softened exactly as is hardened steel. The composition of this alloy usually varies within the following limits: Copper, 3 to 4.5 per cent.; magnesium, 0.4 to 1.0 per cent.; manganese, 0 to 0.7 per cent.; aluminum, balance; iron (as impurities), 0.4 to 1 per cent.; silicon, 0.3 to 0.6 per cent. Its density is about 2.85. It is used only in the forged or rolled condition. This alloy has been produced for some years commercially and is in demand for the fabrication of parts for which both lightness and strength are required, such as for aircraft. Its tensile strength will average 50,000 to 60,000 lb. per. sq. in. (3515 to 4218, kg. per sq. cm.) after appropriate heat treatment, such as that described by Wilm.

Jan 6, 1919

By Albert Wiggin

THE copper-bearing sulphide ores from the mines in Butte, Mont., which are for the most part concentrated at the Boston & Montana duction Works in Great Falls and at the Washoe Reduction Works in Anaconda, contain the following minerals of economic value: Chalcocite, Cu2S; enargite, Cu3AsS4; cupriferous pyrite; bornite, Cu3FeS3; covellite, CUS; tetrahedrite, 4 CuS2Sb2S3; and tennantite, Cu8AsS7;. All of the ores carry a high percentage of pyrite. Many of the above sulphides are argentiferous and also carry from 10 to 20 cts. per ton in gold. The gangue is quartz and highly altered granite, with sphalerite and galena and more rarely barite and hubnerite as accessory minerals. An average chemical analysis of the second-class ore concentrated at Great Falls and Anaconda would be as follows: Copper 3.2 to 3.4 per cent. Silver 2. 3 to 3.0 oz. per ton. Gold 0.0-1 to 0.15 oz. per ton. Si02.. 55 to 60 per cent. FeO 12 to 15 per cent.. Sulphur 12 to 1.5 per cent. Al2O3, S to 10 per cent. CaO 0.5 to 0.8 percent. As203, 0.5 to 0.6 percent. Lead 0.2 to 0.3 percent. Before describing the Great Falls flow sleet as installed, at Anaconda, a brief history of the Great Falls concentrator will be given, including a description of the various concentrating machines and systems which have been tested in the search for a flow sheet which would give a higher recovery of copper.

Jan 8, 1913

By C.F. Pascoe, H.J. Roast

The object of this investigation was to compare the arsenical antimony-lead alloy with some of the regular bearing-metal alloys. With this end in view, the following tests were made: 1. Chemical analyses to show the composition. 2. Thermal analyses. 3. Macroscopic examination of fractures. 4.Microscopic examination. 5. Tensile tests. 6. Crushing tests. 7. Scleroscope tests.' 8. Brinell tests, at temperatures from 80' to 400' F. Tests 3 to 8 were conducted on both sand-cast and chill-cast specimens. For the purposes of this paper the alloys have been numbered from 1 to 6; their composition is as follows: Lead. ' Antimony, Tin, CoppeR ARSEnic, NumBeR PeR CENT. PER CENT. PeR Cent. PeR CeNt. PeR Cent. 1 82.64 12.95 4.40 0.01 none 2 27.54 9.04 59.87 2.65 none I 3 none- 0.13 ------------1.47 4 82.95 16.85 none 0.20 none 5 83.78 15.35 none 0.05 0.82 6 0.15 7.87 . 84.13 7.85 none The analysis, in the case of the antimony, was made by solution in sul-furic acid and titrated with permanganate. For tin, the ferric-chloride method was used; for copper, the alkaline separations, and the thio-cyanate methods; for arsenic, volatilization of arsenic and titration with iodine. The lead, with the exception of alloy 6, was taken by difference. With the exception of No. 5, the alloys were taken as received from the manufacturer. About 40 lb, were melted in a graphite crucible and recast into small ingots to insure uniformity of composition; samples

Jan 1, 1923

By C. F. Pascoe, H. J. Roast

The object of this investigation was to compare the arsenical antimony-lead alloy with some of the regular bearing-metal alloys. With this end in view, the following tests were made: 1. Chemical analyses to show the composition. 2. Thermal analyses. 3. Macroscopic examination of fractures. 4.Microscopic examination. 5. Tensile tests. 6. Crushing tests. 7. Scleroscope tests.' 8. Brinell tests, at temperatures from 80' to 400' F. Tests 3 to 8 were conducted on both sand-cast and chill-cast specimens. For the purposes of this paper the alloys have been numbered from 1 to 6; their composition is as follows: Lead. ' Antimony, Tin, CoppeR ARSEnic, NumBeR PeR CENT. PER CENT. PeR Cent. PeR CeNt. PeR Cent. 1 82.64 12.95 4.40 0.01 none 2 27.54 9.04 59.87 2.65 none I 3 none- 0.13 ------------1.47 4 82.95 16.85 none 0.20 none 5 83.78 15.35 none 0.05 0.82 6 0.15 7.87 . 84.13 7.85 none The analysis, in the case of the antimony, was made by solution in sul-furic acid and titrated with permanganate. For tin, the ferric-chloride method was used; for copper, the alkaline separations, and the thio-cyanate methods; for arsenic, volatilization of arsenic and titration with iodine. The lead, with the exception of alloy 6, was taken by difference. With the exception of No. 5, the alloys were taken as received from the manufacturer. About 40 lb, were melted in a graphite crucible and recast into small ingots to insure uniformity of composition; samples

Jan 1, 1923

By B. R. Palmer, M. C. Fuerstenau

Oxide and silicates respond to flotation with a large number of anionic collectors. These include carboxylates (fatty acids), sulfonates, alkyl sulfates, and certain chelating agents. In contrast to sulfide mineral flotation, relatively long-chained collectors (generally longer than 10 carbons) are required to effect flotation of this group of minerals. The structural formulas of these various anionic collectors are presented in Table 1. Formulas and solubilities of various fatty acids are presented in Table 2. These solubilities are those of the dissolved and undissociated species of the acid, e.g. oleic acid, HOl(aq). Of the compounds listed, oleic acid and abietic acid, which IS a major constituent of talloel, are used most commonly in commercial applications. The usage of these two collectors is related to cost and effectiveness. In addition to their high molecular weights, oleic acid contains a double bond in its hydrocarbon chain, while abietic acid is comprised of aromatic rings containing double bonds. The presence of double bonds increases solubility significantly over that of saturated homologues. Stability of metal-collectors assumes a most significant role in these systems, and the solubility products of various metal carboxylates and sulfonates are listed in Tables 3, 4 and 5. In addition to these properties, strength of the acids of these various anionic collectors is also important. The carboxylic and hydroxamic acids are weak acids whereas sulfonic and alkyl sulfuric acids are strong acids. All of the carboxylic acids fall in the range of pK = 4.7 ± 0.5 ([2]); the pK of hydroxamic acid is around 9 ([3]); and the pK of sulfonic acid is about 1.5 ([4]).

Jan 1, 1976

By Jed H. Mosgrove, Paul M. Budzak

INTRODUCTION Good mine ventilation starts at the drawing board. A coal mine will either be adequately ventilated or not, depending upon how good a job was done in the planning. Poor planning will not only result in inadequate ventilation, but may also lead to low productivity. A mine foreman cannot achieve high productivity while trying to maintain 4.2 m3/s (9000 cfm) of air at his last open crosscut. PLANNING What will provide adequate ventilation? A rule of thumb in laying out a coal mine is to provide enough entries in the main entry system to keep air velocities at 2.5 m/s (500 fpm) or less; at this rate of flow, good ventilation may be expected. The intake and return air courses of a coal mine do not improve with age. Entries will deteriorate, ribs will slough, roof will fall, and clutter will accumulate. The resistance offered to air travel will continue to increase with each passing year. Providing enough airways for the present and the future cannot be emphasized too strongly. In order to better appreciate good planning and bring it into better focus, assume a hypothetical mine has the following features: entries 1.5 m (5 ft) high and 6.1 m (20 ft) wide; one intake and one return entry; the mine development is 4.8 km (3 miles) from the intake and upcast ventilation shafts; and the ventilation requirement is 141.6 m3/s (300,000 cfm). Air travel of 4.8 km (3 miles) into a mine and the same distance back to the fan is not unusual. However, to do so with one intake and one

Jan 1, 1981

By F. Louis Grammer

THE difference between blast-furnace gas and ordinary producer-gas is chiefly that blast-furnace gas is richer in CO, and poorer in hydrocarbons and hydrogen, as is shown in Table I. TABLE I.-Volume-Composition of Producer-Gas and Blast-Furnace Gas.. Producer-Gas. Blast-Furnace Gas. Per Cent. Per Cent. (Volume.) (Volume.) CO . 22 to 33 25.5 00., . 6 to 3 13.0 H . . 13 to 7 1.5 CH, 3 to 1.5 0.0 N . 54 to 60 60.0 If natural gas, water-gas, or by-product coke-oven gas be available, blast-furnace gas can be made to resemble very closely in calorific value the average producer-gas by merely mixing in one of these products, which are richer in hydro-carbons and, hydrogen. Blast-furnace gas, however, possesses a virtue which is sometimes absent in producer-gas, especially with coals very high in sulphur-namely, that blast-furnace gas is relatively free from sulphur. In some regions the sulphur in coal is an acute problem, and it is highly desirable to find uses for the products which are specially high in sulphur-content. An advantage would be gained if coal was coked before charging into the gas-producers, but this adds to the expense. Sulphur may be partly eliminated from coal by washing it, the success depending upon the form in which the sulphur occurs, and upon the efficiency of the washer, but it cannot always be removed to the extent desired at most American plants. It is not practicable to remove all the sulphur from producer-gas by washing it, though Mr. Reinhardt, of Germany,

Sep 1, 1908

By Jed H. Mosgrove, Paul M. Budzak

INTRODUCTION Good mine ventilation starts at the drawing board. A coal mine will either be adequately ventilated or not, depending upon how good a job was done in the planning. Poor planning will not only result in inadequate ventilation, but may also lead to low productivity. A mine foreman cannot achieve high productivity while trying to maintain 4.2 m3/s (9000 cfm) of air at his last open crosscut. PLANNING What will provide adequate ventilation? A rule of thumb in laying out a coal mine is to provide enough entries in the main entry system to keep air velocities at 2.5 m/s (500 fpm) or less; at this rate of flow, good ventilation may be expected. The intake and return air courses of a coal mine do not improve with age. Entries will deteriorate, ribs will slough, roof will fall, and clutter will accumulate. The resistance offered to air travel will continue to increase with each passing year. Providing enough airways for the present and the future cannot be emphasized too strongly. In order to better appreciate good planning and bring it into better focus, assume a hypothetical mine has the following features: entries 1.5 m (5 ft) high and 6.1 m (20 ft) wide; one intake and one return entry; the mine development is 4.8 km (3 miles) from the intake and upcast ventilation shafts; and the ventilation requirement is 141.6 m3/s (300,000 cfm) . Air travel of 4.8 km (3 miles) into a mine and the same distance back to the fan is not unusual. However, to do so with one intake and one

Jan 1, 1981

By W. F. Holbrock

IN a previous paper1 a method for the measurement of the compara-tive desulphurizing power of slags was described and data, were presented covering the range of likely slags containing up to 10 per cent MgO. In the present work the range is extended to include Tags containing up to 20 per cent MgO, and the effect of temperature on 5 per cent and 15 per cent MgO sings over the range 147.5° to 1525° C. is included. The method used was that described in the previous paper except that some of the temperature measurements were made with a noble-metal thermocouple instead of a tungsten-graphite thermocouple. Twelve-gram charges of 0.S per cent sulphur-metal and 5-gram charges of slag were subjected to heating for 30 min. at a prescribed temperature. The charges were analyzed for sulphur content. For purposes of comparison, desulphuriz-ing power is defined as the ratio of the percentage of sulphur found in the slag to that found in the metal. Six slag and metal charges were made coincidentally in holes in a section of all electrode graphite rod. One slag and metal charge was used as a control by means of which adjustment in the desulphurization value could be made for small varia-tions in temperature and other factors. For convenient presentation of the data, the slags are arranged in groups -in each of which one of the four main slag components is constant. Thus, the 15 per cent MgO slags are presented separately from those containing 20 per cent MgO.

Jan 1, 1938

By Fred W. Stiefel

To open the Fisher Hill mine of the Republic Steel Corporation, it was necessary to sink an inclined shaft into the rock and excavate stations, drifts, and ore pockets. This inclined shaft, or slope, q ft. 8 in. high by 28 ft. 2 in. wide, has compartments for two 5-ft. gauge ore-skip tracks of 120-lb. rail, one 5-ft. gauge man-skip track of 60-lb. rail, a concrete stairway, and service pipes consisting of a 12-in. compressed-air line, an 8-in. pump-discharge line, and a 4-in. water-supply line. When we were requested to investigate the work, the progress being made (by another organization) amounted to only 2 ½ ft. advance in 24 hr. We reported that progress of at least 6 ft. per day could be made. As a result, the Defense Plant Corporation, through the Republic Steel [ ] Corporation, entered into a contract with us to sink the shaft. The contract contained a penalty and bonus clause predicated upon an advance of 150 ft. per month. The average advance was 7.8 ft. per day; that in the best six-day week was 60 ft. and that in the best month, 215.5 ft. in 25 working days. The drilling was performed by 3 ½ -in. wet drifter drills mounted on five 8-ft. columns (see accompanying illustration). Even though a full-face drilling and mucking operation is customarily desirable, the steep slope, with the accumulation of water and fine debris at the face, made a modified heading and bench operation preferable. The heading and bench were of equal height. The five columns, instead

Jan 1, 1945

By F. Louis Grammer

The difference between blast-furnace gas and ordinary producer-gas is chiefly that blast-furnace ga,s is higher in CO2 and lower in hydrocprbons and hydrogen, as is shown in Table I. Table I.— Volume-Composition of Producer-Gas and Blast-Furnace Gas. Producer-Gas. Blast-Furnace Gas. Per Cent. Per Cent. (Volume.) (Volume.) CO......22 to 33 25.5 CO2........ 6 to 3 13.0 H...13 to 7 1.5 CH4........ 3 to 1.5 0.0 N..54 to 60 60.0 If natural gas, water-gas, or by-product coke-oven gas be available, blast-furnace gas can be made to resemble very closely in calorific value the average producer-gas by merely mixing in one of these products, which are richer in hydrocarbons and hydrogen. Blast-furnace gas, however, possesses a virtue which is sometimes absent in producer-gas, especially with coals very high in sulphur—namely, that blast-furnace gas is relatively free from sulphur. In some regions the sulphur in coal is an acute problem, and it is highly desirable to find uses for the mine-products which are specially high in sulphur-content. An advantage would be gained if the coal were coked before charging into the gas-producers, but this adds to the expense. Sulphur may be partly eliminated from coal by washing it, the success depending upon the form in which the sulphur occurs, and upon the efficiency of the washer, but it cannot always be removed to the extent desired at most American plants. It is not practicable to remove all the sulphur from producer-gas by washing it, though Mr. Reinhardt, of Germany,

Jan 1, 1909

By Anthony Sharkey

Introduction In Europe, longwall mining accounts for over 95% of coal produced from deep mines, whereas in the U.S.A., longwall accounts for less than 5% of deep mine coal production. There are many reasons for the difference, but in general: a. U.S. mine operators have invested heavily in continuous mining or conventional equipment which continues to produce acceptable results; b. There 1s a reluctance to change from, a familiar (and proven) mining method to an unfamiliar one, despite promises of huge production increases,; c. Multi-seam mining in Europe has made longwall the only really viable mining method, whereas In the U.S.A. multi-seam mining (particularly in "below-drainage" mines), is relatively rare; d. Although some spectacular results have been achieved by longwall in the U.S.A.(I), the general performance of longwall in this country has not been good enough to generate much enthusiasm for its wide application; e. The much higher level of investment in face equipment required for longwall when compared to alternative methods (approximately $7.5 million for a medium seam longwall installation and $1 million to $1.5 million for a continuous miner section). However, declining productivity, national energy requirements and safety legislation are creating a situation in which longwall as a viable mining method is being examined more closely and in more places than hitherto. This paper, therefore, attempts to identify some of the principal criteria needing examination in any assessment of the feasibility of longwall application, the limitations of information for those criteria in the decision making process and the need for further research and development to assist the mine operators make the decisions.

Jan 1, 1981