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By W. H. Pailes, V. Castranova, K. Van Dyke, J. H. Kang

Recruitment and activation of phagocytic cells after inhalation of occupational dusts has been implicated in the development of pulmonary inflammation, fibrosis, or emphysema. Plateletactivating factor has been described recently as a potent inflammatory agent. The objective of the present study was to investigate if platelet-activating factor might play a role in the inflammatory response of the lung to occupational dusts. In vitro exposure of alveolar macrophages to cotton dust (4 mg/ml), silica (10 mg/ml), or coal dust (10 mg/ml) results in the release of picomole levels of platelet-activating factor with a potency of cotton dust > silica > coal dust. Time course studies indicate that platelet-activating factor release from alveolar macrophages is maximal approximately 10 minutes after in vitro exposure to silica. In vitro, platelet-activating factor is a direct stimulant of polymorphonuclear leukocytes increasing chemiluminescence 2.3 ± 0.3 fold above control. The maximal response occurs in the dose range of 10-10 to 10-7M platelet activating factor. In addition, plateletactivating factor potentiates the response of polymorphonuclear leukocytes to a second stimulant; i.e., chemiluminescence generated from polymorphonuclear leukocytes exposed to plateletactivating factor (10-9M) plus chemotactic peptide (FMLP, 10-9M) is 88 ± 22% greater than the sum of the separate effects of these stimulants. In contrast, platelet-activating factor is not a direct stimulant of alveolar macrophages in vitro, having no stimulatory effect on chemiluminescence or superoxide anion release even at levels of 10-6 to 10-5M platelet-activating factor. However, platelet-activating factor (10-5M) does potentiate zymosan-stimulated activation of chemiluminescence and superoxide release from alveolar macrophages by 117 ± 56% and 30 ± 4%, respectively. In conclusion, these data indicate that exposure of the lung to occupational dusts results in the release of platelet-activating factor which may activate polymorphonuclear leukocytes and potentiate the responsiveness of alveolar macrophages to dust. The net result would be pulmonary inflammation and hypersecretion of reactive species by pulmonary phagocytes which could lead to tissue damage and pulmonary disease.

Jan 1, 1991

By W. M. Small

From the one producing well in Austria, 7000 bbl. of heavy oil was produced and sold in 1933. This well, owned by the Raky-Danubia, was completed late in 1932 in the formation called Flysch (Upper Cretaceous or lower Eocene) near Zistersdorf north of Vienna. Two wildcats drilling in the vicinity have not yet reached the producing beds. Near Vienna, the European Gas and Electric Co. drilled five shallow wells for gas, three of which were failures. One gas well was completed for 8 million cu. ft. capacity at 260 meters depth and with a pressure of 570 lb. Another is being tested following blowouts. One gas well and one dry hole were completed in the pool in 1932. In Upper Austria, the same company abandoned one wildcat at 679 m. and is drilling at 280 m. on a second. No gas was sold during the year, but a line is being completed to supply fuel to boilers of the Vienna Electric Works. Hungary No deep drilling was done in 1933. Oil and gas concession was granted by Parliament to the European Gas and Electric Co. on Trans-danubia. Active geological and geophysical investigation is in progress. Czechoslovakia In the Gbely and Hodonin shallow fields, some 26 wells were drilled of which three were reported as failures. The total production of the fields was 111,736 bbl. of oil and 40 million cu. ft. of gas used as fuel in the field. At Vacenovice, where there were already two small gas wells, a third producer was completed at 420 m. This pool markets a small amount of gas, which is compressed and used to light railway carriages. Wildcat wells were reported drilling at about 300 m. depth at both Jasina and Telnice. One 700-ni. wildcat near the Hodonin field was abandoned. Another failure was reported completed and abandoned at 820 m. near Holic.

Jan 1, 1934

By Eero Pokki

The Päijänne Tunnel is being constructed to conduct raw water from Lake Päijänne to the Helsinki Metropolitan Area. The length of the tunnel will be 120 km and the cross-sectional area is 15.5 sq m. Excavation started in 1973 and will be completed in 1982. 94 km of the tunnel have been constructed from Lake Päijänne towards Helsinki. The construction has advanced according to plan, and due to Finland's hard bedrock there have not been any great surprises. When completed, the tunnel will meet the water requirements of the Helsinki Metropolitan Area for a long time in the future because, with a free fall of 36 m, 9.6 cu m/s of water can be conducted through the tunnel from Lake Päijänne to the Helsinki Area. With pumping it will be possible to obtain a flow of 15 cu m/s, although at first the discharge will only be 5 cu m/s. The tunnel was partly taken into service as early as March 1979 when it was possible to conduct water to the municipalities north of the Helsinki Area. It is already possible now to conduct water to Helsinki, too, but from the Korpimäki pumping station onwards it must be done through the River Vantaa, which is polluted. When the tunnel is completed in 1982, good quality water can be conducted from Lake Päijänne to the Helsinki Metropolitan Area through a continuous rock tunnel. The cost estimate in 1973 was $ 88 million while it is now $ 90 million notwithstanding high inflation. AREAS NEEDED FOR CONSTRUCTION The areas needed for construction were bought through voluntary deals. Areas of 4 - 5 hectares were bought along the tunnel line at about 5 km intervals. The excavation of access tunnels was started from these places, and the crushed rock from the main tunnel was also transported there. However, it was not necessary to buy the area

Jan 1, 1979

By G. H. Clamer

Page Bronze Bearing Metals. By G. H. Clamer............... 162 Pennsylvania Railroad Anti-friction and Bell Metals. By F. M. Waring .. 166 The Tin-plate Industry. By D. M. BUck. Discussed by G. H. ClameR, J. W. RICHARDS............................ 168 The Aluminum Bronze Industry. By W. M. Corse............ 171 Bronzes, Bearing Metals, and Solders. By G. K. BURgess and R. W. Woodward. Discussed by G. H. Clamer, R. T. Roberts......... 175 Babbitts and Solder. By G. W. Thompson................ 184 The Cadmium Supply of the United States. By C. E. Siebenthal. Discussed by M. L. Lissberger, C. W. Hill, F. F. ColcoRd.......... 185 Solder, Its Use and Abuse. By M. L. LissbergeR............ 192 Bronze Bearing Metals G. H. ClameR,* Philadelphia, Pa.—Unfortunately, prior to the war no serious attention was given to the conservation of tin, notwithstanding that this country is practically dependent upon outside sources for its entire tin supply. Tin is of vital importance in many industries, but it is surprising how many and how excellent are its substitutes when we become acquainted with them. Tin has always been a relatively high-priced metal, and it is part of the human attitude to associate high prices with high standards; it is not until the price of a commodity becomes well-nigh prohibitive that we hunt for substitutes, because the idea of substitution seems always to involve an assumption that the substitute must necessarily be inferior. The history of the development of bearing bronzes is a striking example of this policy. In the early days, copper-tin alloys were almost universally used, the idea then being prevalent, which is still held by many, that a bearing to resist wear must be hard, and the harder the better. The favorite bronze bearing contained 90 per cent. copper and 10 per cent. tin; frequently, in service which was considered severe, even higher proportions of tin were used. Such hard alloys have great resistance to compression, but as a rule they had a very wide factor of safety in this

Jan 1, 1919

By Douglas A. Karvonen

INTRODUCTION Although the economy in the Province of Saskatchewan has historically been agriculturally oriented, a major source of wealth has been realized through natural resources such as petroleum and natural gas, coal and uranium and since the 1960's - potash. Potash or potassium chloride, in the form in which it is presently produced in Saskatchewan, is used primarily as fertilizer, a source of potassium, one of three essential elements required for healthy plant growth, with about 5% of production sold to the chemical industry for use in such products as soap and liquid detergent, glass and ceramics, textiles and some pharmaceuticals. Saskatchewan is estimated to have approximately 40% of known world potash reserves and its ten active mines account for some 25% of annual world production. POTASH CORPORATION OF SASKATCHEWAN In an effort to guarantee a fair and lasting economic benefit to the Province of Saskatchewan, the Potash Corporation of Saskatchewan (PCS) was created on February 4, 1975 through an Order-in-Council under the Crown Corporation's Act and later on April 1, 1976, the Potash Corporation Act provided for its continuation. PCS was given the mandate to acquire ownership of a significant portion of the potash mining capacity in Saskatchewan and to establish an organization capable of operating the facilities, marketing the product, and conducting all of the other activities necessary for a successful corporation competing in the world potash industry. The corporation is the largest potash producing and selling organization in the Western World - or third largest in the world. (See also Figure 1). PCS through its wholly-owned subsidiary, PCS Mining, has acquired a whole or partial interest in five geographically distinct producing properties or divisions, representing today more than 40% of Saskatchewan's potash producing capacity. The acquisition of each division was followed by a program designed to exploit and maximize the particular opportunities at each minesite for increased production. To that end, PCS Mining has taken an active lead in expanding to meet future demand and has implemented expansion programs at all properties except for its Esterhazy Division where it has a long-term processing agreement with International Minerals & Chemical Corporation (Canada) Limited. Through its wholly-owned subsidiaries, PCS Mining, PCS Sales and PCS Transport, this provincial crown corporation has the ability to operate as a producer and marketer of Saskatchewan's potash. In certain instances PCS, the parent company, assists in the financing of various projects undertaken by the subsidiaries and in other cases the subsidiary involved arranges its own financing. POTASH RESERVES As a background to the specific issue of financing, it is helpful to describe briefly the nature of the deposit which supports the industry - our collateral, if you like. The potash beds underlying Saskatchewan are sedimentary rock deposits formed by the gradual evaporation of an inland sea during the Middle Devonian period roughly 400 million years ago. Now known as the Prairie Evaporite, this formation stretches from North Dakota and Northeast Montana through Southern Saskatchewan into North-Central Alberta. These beds lie at mining depths ranging from 950 - 1000m (3,000 ft) below the surface, where conventional mining methods can be used, and to 1,650m (5,000 ft), where solution mining must be

Jan 1, 1985

By Y. Sagawa

The Hishikari Gold Mine consists of epithermal vein type Au-Ag deposits, each of which has several veins standing almost in parallel each other and inclining 70 to 90 degrees. The veins have 1 to 3m widths and are extracted mainly by drifting and bench stoping with waste rock fill. The Hishikari has tried to optimise the stope dimensions by the Potvin’s Stability Graph Method, which is well known as an easy and effective tool to design appropriate stope dimensions. The method, however, has some limitations of application due to its simplicity so that it has been used only in limited cases. The authors developed a new numerical approach, based on 2D FEM solu-tion for the stability analysis. As a result, it has been possible to solve the Hishikari situation. The authors also introduce the concept of the design approach and compare the results obtained by us-ing the approach with the database of the Stability Graph Method for validation. La mine d’or Hishikari se compose de gisements d'Au-AG avec des types de filons épither-maux dont chacun a plusieurs veines qui sont presque parallèle les unes par rapport aux autres et sont inclinées de 70 à 90 degrés. Les veines ont de 1 à 3 m de largueur et elles sont principalement extraites en creusant et en abattant des bancs de roches encaissantes. Hishikari a essayé d’optimiser les dimensions du chantier d’exploitation avec la méthode de graphique de stabilité de Potvin, bien connue comme un outil facile et efficace pour avoir des dimensions de chantier d’exploitation appropriées. Toutefois, la méthode a certaines limites d’application dues à sa sim-plicité de sorte qu’elle a été utilisée seulement dans des cas limités. Les auteurs ont développé une nouvelle approche numérique, basée sur la solution 2D FEM, pour l’analyse de stabilité. Par conséquent, il a été possible d'expliquer la situation de Hishikari. Les auteurs présentent le concept de l'approche et ils comparent les résultats donnés en l'employant avec la base de données de la méthode de graphique de stabilité pour la validation. Die Hishikari-Goldmine besteht aus Au-Ag-Ablagerungen mit epithermalen Adern, die jeweils mehrere Adern aufweisen, die mit einer Neigung zwischen 70 und 90 Grad nahezu parallel zueinander verlaufen. Die Adern haben eine Breite von 1 bis 3 Metern und werden überwiegend durch Vortrieb von Stollen und Bergversatz abgebaut. Hishikari hat versucht, die Abmessungen des Abbauorts durch das Potvins Stabilitätskurvenverfahren zu optimieren, das als einfaches und effizientes Mittel für die Erzielung angemessener Abmessungen des Abbauorts bekannt ist. Das Verfahren weist jedoch aufgrund seiner Einfachheit einige Anwendungsbeschränkungen auf, so dass es nur in einigen Fällen benutzt worden ist. Vor diesem Hintergrund haben die Autoren einen neuen numerischen Ansatz entwickelt, der hinsichtlich der Stabilitätsanalyse auf der 2D-FEM-Lösung beruht. Dadurch wurde es möglich, das Problem von Hishikari zu lösen. In diesem Dokument stellen die Autoren die Grundlagen des Ansatzes vor und vergleichen die erzielten Resultate durch Benutzung der Datenbank des Stabilitätskurvenverfahrens zur Validierung.

Jan 1, 2003

[NORTH AMERICA UNITED STATES ALABAMA ADAMSVILLE U. S. Steel Corp. Neason, James E. ALBERTVILLE Thompson Floral Co. Loudermilk, E. L. ANNISTON No Data Supplied Bonnichsen, Bill BESSEMER U. S. Steel Corp.-TCI Div. Beck, Adolph W., Jr. Boyle, James R. Cofield, William H. *Evans, James D. F., II *Guyton, Jesse C. Johnson, Leland H. Tetley, Albert L. No Data Supplied Crockard, F. H. Fisher, Joe T. Griffin, Joseph M. McFarland, E. A. Mohr, Henry E. Neal, George M. Walker, Wilber D.]

Jan 1, 1961

By AIME AIME

AWIDE variety of subjects was discussed at the Wednesday sessions on Non-metallic Minerals. W. M. Weigel as chairman, presided at the morning session, and W. M. Myers, vice-chairman, in the afternoon. In the morning session a paper by W. N. Logan on "Mineral Wool Resources of Indiana" was read by H. H. Hughes. Mr. Logan explained that as Indiana is the principal source of mineral wool, the paper might well he entitled "Mineral Wool Resources of the United States."

Jan 1, 1932

By William H. Meadows

This information describes the ventilating and methane gas testing procedures utilized during the raising of a mine shaft using a Robbins 81-R drill. An overview of the raise boring process is presented to illustrate the ventilating and gas testing procedures. A mine shaft project, 4.88 m (16 ft) in diameter and approximately 457.2 m (1500 ft) in length, will be the focal point of the presentation, with reference to similar projects for additional comments on variations in procedure.

Jan 1, 1980

ASSOCIATED MINERS, UNIVERSITY OF IDAHO The following officers of the Associated Miners of the University of Idaho have been elected for the academic year for 1918-1919: President, Lloyd A. McDougal; vice-president, Russel D. Bowers; secretary-treasurer, Charles M. Otter. While the society has been somewhat quiescent during the existence of the S. A. T. C., we are now looking forward to an active session with the resumption of normal university life. CHARLES M. OTTER, Sec'y-Treas.

Jan 2, 1919

By Will D. P

A 328 m, 2.4 m diameter raise-bored shaft has been equipped with a modular ladderway system to provide a second means of egress from Aberfoyle's Hellyer Mine in Tasmania. The modular ladderway consists of platforms, ladders and service pipes assembled at the surface and progressively lowered down the shaft. This system provides a rapid, safe and economical non-entry approach to equipping a shaft. The modular construction and single suspension allows for future extension or for reuse in another shaft.

Jan 1, 1988

[HORACE MOSES President Gamerco, New Mexico BENEDICT SHUBART Secretary-Treasurer 521 Boston Bldg., Denver, Colorado ROBT. L. HAIR Vice-President for Colorado VIRGIL M. McKNIGHT Vice-President for New Mexico A. E. GIBSON Vice-President for Utah THOS. FOSTER Vice-President for Wyoming EXECUTIVE BOARD COLORADO: J. B. DICK Walsenburg C. M. BILLINGTON Lafayette NEW MEXICO: W. C. HOLMAN Dawson ROBT. WYPER Gallup UTAH: R. R. KIRKPATRICK Sunnyside GEO. A. SCHULTZ Latuda Wyoming: EDWARD BOTTOMLEY Kleenburn QUINCE RICE Superior]

Jan 1, 1926

By Brad Thrall

"Viceroy Resource Corporation of Vancouver, B.C. has recently developed the Brewery Creek Mine, an open pit heap leach gold mine 55km East of Dawson City in Yukon, Canada. Mineable reserves are 17.1 million tonnes of ore grading 1.45 grams/tonne, containing 24.8 tonnes (797000oz) of gold. Ore processing employs the heap leach technology using run of mine ore. 1996 highlights included construction of 158,000 m2 of leach pad, placing 1.9 million tonnes of ore at a grade of 1.46 gpt on to the leach pad, construction and successful commissioning of the ADR plant, and pouring of the first gold bar on November 15th.Permitting this mine has drawn significant attention because the heap leach technology, while proven in other jurisdictions, has not been used in Yukon where winter site temperatures routinely reach -45 C for extended periods.IntroductionFollowing a comprehensive governmental and public review of the proposed project, a Class A Yukon Water license was issued in August, 1995. Construction started in 1995 and leaching commenced in October 1996. The Brewery Creek Mine is in Yukon, the home of the Klondike Gold Rush nearly 100 years ago. This is the only part of Canada not glaciated in the Wisconsin advance so there is oxide ore resulting from about 80,000 years of weathering -unique in Canada. Discontinuous permafrost is present on site and has required special consideration in the design. Winter temperatures dip below minus 50 C, demanding an approach to mining and leaching that is different than in more temperate climates.Project DescriptionThe mine comprises 8 shallow, largely oxidized gold deposits along a 7km corridor. An 8 year mining program will extract 17.1 million tonnes of ore grading 1.45 grams/tonne, containing 24.8 tonnes (797,000 oz) of gold. The strip ratio will be 1.35: 1. Overall gold recovery is projected at 78%. There will be 2.3 million tonnes of ore placed on the pad annually, with an average of 2.57 tonnes (83, 000 oz) of gold recovered per year. Because of the cold winters, the mine will operate approximately 210 days per year, while the leaching process will operate on a year round basis. Run of mine ore is placed on the heap 150 days per year. The recovery process utilizes conventional sodium cyanide leaching , carbon adsorption, caustic desorption, and on-site smelting. Peak employment during mining is 145 and winter staff is reduced to 45."

Jan 1, 1997

By Gerald Edmunds

In a paper1 before this Institute in 1940, the writer reported that the surface orientation texture of zinc and cadmium differed from the texture existing within the casting, in that basal planes were parallel to the surface. He was unable to rationalize the , origin of this surface texture. It is conceivable that the surface orientation texture is influenced by the atomic force fields, which are different at the surface and within the body of metal. It was considered that a radical changc in the surface against which the metal is cast might alter the surface texture. In trial of this, zinc has been cast against a surface of molten lead, but the same result has been obtained—i.e., (001) parallel, ± 2oo, to the surface—as was obtained when it was cast against air, Transite or iron.' Into the bottom of an open cast-iron tray, 7 by 14 in. by 2 in. deep, preheated to about 275°C., about 1/4 in. of lead, then 1/2 in. of zinc was poured, and the whole was covered with an electric heater to enforce freezing from the bottom upward. After a lapse of time sufficient for the zinc but not the lead to solidify, the zinc slab was raised free of the molten lead. The bottom surface of the zinc casting was coated with a thin layer of lead. A specimen of the slab 136 in. square was metallographically polished to flatten it and remove the slight film of lead, and the orientation was determined by the method used before,' with the result stated above. Some determinations have also been made on other metals, as follows: Surface orientation determinations were made on the 3/8-in. diameter grip section of two 1/4-in. diameter aluminum-alloy die-cast tensile bars and on a 1/8 by 1/2-in. flat aluminum-alloy die-cast tensile bar. None of these three aluminum die castings exhibited any preferred orientation at the surface. A brass die casting (nominal composition: 60 Cu, 37.75 Zn, 1.00 Sn, 1.00 Pb, 0.10 Al, 0.15 Si) furnished by C. S. Smith was also examined. The surface showed a random orientation of alpha and beta. Summary A zinc casting solidified against a molten lead surface was found to have the same surface orientation texture, (001) parallel to the surface, as other zinc and cadmium castings. Aluminum and alpha-beta brass die castings were found to have random grain orientation textures at the surface. Acknowledgment The X-ray determinations were conducted by M. L. Fuller.

Jan 1, 1945

By E. O. Mixon, Joseph P. Sehulster

This paper describes the driving of a 1421 m long by 2.74 m diameter sewer tunnel in postglacial soils of the Menomonee River Valley in Milwaukee, Wisconsin. An earth pressure balance machine was used for excavation and to install a "one pass" pre-cast concrete gasketed segmental liner. Also discussed is the use of ground freezing to stabilize the ground for the start of tunneling operations and modifications made to two structures to accommodate final tie-ins for the sewer tunnel.

Jan 1, 1989

By C. D. Breeds, F. E. Valencia, J. H. Pye

The "FIRST AUTOMATIC SHOTCRETE TECHNIQUE" (F.A.S.T.) system, provides safe, efficient methods of remotely lining, shafts/raises. Two systems are available; one system will be used to shotcrete a 3.04m (10-ft.) diameter ventilation raise at a mine in Mexico; the other system incorporating remote control and sensing devices will to used to shotcrete ventilation raises from 2.75-4.9m (9-16 ft.) in diameter. This paper describes development of the system, from idea to field use.

Jan 1, 1981

By R. T. Hukki

This paper presents a series of equations for mechanical ball wear, relating parameters of ball size, mill speed, and mill diameter. The fundamental equation, Eq. 12, presented here is introduced to correlate these basic parameters and thus define and clarify the concept of ball wear. This equation is offered as a general rule, which may be modified to apply to individual problems of grinding. BALL wear as observed in grinding installations is the combined result of mechanical wear and corrosion. Corrosion should be a linear function of the ball surface available. Ball corrosion, however, has been studied so little that its effect, although of great importance, cannot be included in the analyses given here. In a separate paper' it is shown that 1 n = 0.7663 np----=— rpm [1] vD P = c, np D kw [2] T=c²(np)n De tph [3] In these equations n — actual mill speed, rpm np = calculated percentage critical speed D = ID of mill in feet P = power required to operate a mill, kw T = capacity of a mill, tph C¹ and c² - appropriate constants in = exponent of numerical value of 1 5 m 1.5 Exponent m is the slope of a straight line on logarithmic paper relating mill speed (on the abscissa) and mill capacity (on the ordinate). It is generally accepted, although not sharply defined, that ball wear in mills running at low (cascading) speeds is a function of the ball surface available. Accordingly, the wear of a single ball may be considered to be a homogeneous, linear function of its surface and of the distance traveled. Thus dw = f¹(d2) . f2(ds) [4] where dw is the wear of a single ball in time dt, d the diameter of the average ball in ball charge, and ds the distance traveled by the ball in time dt. Indicating that ds - a D n dt, the wear of the average ball in time dt becomes dw = f¹(d2) . f2(Dn dt) 1 --- f¹ (d1) f² (D c3 np-----— dt) \/D = c,d² n, D dt The rate of wear of the average ball is given by dw/dt. dw/dt = c, d² np D lb per hr [5] The weight of the ball charge per unit of mill length is a function of D The number of balls of size d in the ball charge is = f³(D2)/f4(d³). The rate of wear of the total ball charge equals the number of balls times rate of wear of the average ball. Thus rate of total ball wear = — . (dw/dt) w. c, . (l/d) . n,, D lb per hr [6] which is the equation of ball wear in low speed mills. In a mill running at a low speed, grinding is the result of rubbing action within the ball mass and between the ball mass and mill liners. When the speed of the mill is gradually increased toward the critical, the impacting effect of freely falling balls becomes increasingly prominent in comparison with the rubbing action. Reduction of ore takes place partly by rubbing, partly by impact. The share of the freely falling balls in the reduction of ore reaches its practical maximum at a speed somewhat less than the critical; at that speed grinding by rubbing has decreased to a low value. It may be reasonable to think that size reduction by freely falling balls should reach its theoretical maximum at the critical speed, if the fall of the balls were not hindered by the shell of the mill beyond the top point; grinding by rubbing would cease at the critical speed. As a first approximation, wear of freely falling balls may be considered to be a homogeneous, linear function of the force at which they strike pieces of rock and other balls at the toe of the ball charge. The force equals mass times acceleration. The mass of a ball is a function of d3 and its acceleration is a function of the peripheral speed of the mill. The wear of a single ball of size d representing the average ball in a ball charge will therefore be w¹ = f3(F) = f (d3) f7 (v). [7] Indicating that v = D n, and n = c³ np 1/vD, Eq. 7 becomes W1 = cn d3 np Do.5 lb per hr. [8] Total wear of the ball charge equals number of balls times the wear of the average ball. Number of

Jan 1, 1955

A Directory of Computer Programs Applicable to US Mining Practices and Problems. A. L. Sanford, T. L. Myers, and J. F. Stiehr. USBM OFR 131-78, Nov. 1977. A Master Environmental Control and Mine System Design Simulator for Underground Coal Mining. Tape MDS-5. Pennsylvania State University. USBM CT 1.77, Dec. 1975. Contains computer programs on a dynamic general purpose computer simulation model for underground coal mining. A Method for Computing Stabilization Pressures for Excavations in Incompetent Rock. J. D. Dixon and M. A. Mahtab. USBM RI 8128, 1976. Describes technique for determining the confining pressures that must be provided for stabilizing underground openings in incompetent rock. Accident Cost Indicator Model to Estimate Costs to Industry and Society From Work-Related Injuries and Deaths in Underground Coal Mining. Vol. I.III. D. G. DiCanio, A. H. Nakata, D. Colvert, and E. G. LaVeque. OFR 39(1)(3)-77, Dec. 1976. Describes a computer-based model for estimating the tangible costs of injuries and deaths from work-related accidents In underground coal mines. An Interactive Computer System for Evaluating Coal Mine Illumination. R. Goldstein. USBM OFR 110-80, March 1980. Computer system for calculating the illumination on coal mine surfaces due to machine-mounted lights. Analytical Modeling of Mine Roof Behavior Using Statistical Material Properties. R. T. Langland and F. J. Tokarz. USBM UCRL-51876, Aug. 1975. Analytical study of coal mine roof behavior using a computer model of a room-and-pillar mining method. Application of a Total System Surface Mine Simulator to Coal Stripping. Vol. I-VII. Pennsylvania State University. USBM OFR 33(1)(7)-78, Aug. 1978. Describes model development, input and output documentation, three case studies of an open-pit materials handling simulator, and development of two equipment selection models and a cost model. Characterization of the Structural Behavior of Rock Masses. Vol. I-II. L. R. Herrmann and M. A. Taylor. USBM OFR 67(1X2)-75, Sept. 1974. Computer subroutines for characterization of rock behavior, Including variability of properties within the rock mass, planes of weakness orthotropy, strength criteria, and post-failure behavior. Coal Mine Electrical System Evaluation. Vol. I-VII, L. A. Morley, et al. USBM OFF! 61(1)(7)-78, 1977. A concept to improve underground coal mine electrical system safety and availability. This proposed technique is based upon the ability to predict incipient failures in the mine power system. Prediction is made possible by a new method which uses a minicomputer to implement pattern recognition algorithms. Computer Modeling of Fluid Flow During Production and Environmental Restoration Phases of In Situ Uranium Leaching. R. D. Schmidt. USBM RI 8479, 1980. Describes development and application of a computer model for simulating the hydrological activity associated with in situ leaching. Interpretation of Rock Mechanics Data, Vol. II (A Guide to Using UTAH2). W. G. Pariseau. USBM OFR 47(2)-80, June 1978. Brief description of a two-dimensional elastic-plastic finite element program intended for mine stability analysis. Iterative Approximation Techniques for Microseismic Source Location. G. H. Dechman and Meng-Cherng Sun. USBM RI 8254, 1977. Iterative approximation method for microseismic source location developed to adapt microseismic techniques for rock structure analysis to investigations of surface problems such as slope failure in an open-pit mine. Limits and Cost Sensitivity of Alternative Parting Handling Methods. Vol. I-II. T. E. Finch, D. R. Haley, and C. J. Speaks, Jr. USBM OFF! 117(1)(2)-77, March 1977. Programs to examine alternative methods of handling the parting between two coal seams in a surface mining operation. Alternative methods consist of equipment implementation in either a stowing or haulback technique. Methane-Water Flow in Coalbeds. H. S. Price. USBM CT 1-79, April 1972. This computer program calculates transient, two-dimensional, gas¬water flow in a porous medium. Plastic Canopy-A Computer Program for the Structural Analysis of Protective Canopies. K. D. Winters, G. A. Gavan, and J. C. Ault. USBM IC 8795, 1979. Presents a FORTRAN IV program for the structural analysis of protective canopies used In underground mines and rollover protective structures used in surface mines. The Three-Dimensional Structural Analysis of Double-Entry and Single-Entry Coal Mines. Vol. I-III. Agbabia Associates. USBM OFF 16(1)(3)-80, Oct. 1978. Describes the computer program designed to perform a three-dimensional finite element analysis of an instrumented section of double- and single-entry coal mines. Underground Mine Evacuation Plans Analyzed by Computers. D. Tesarik, D. Nicholson, and A. B. Boghani. MSHA, Vol. 4, No. 5, Oct. 1979. USBM has developed a computer program to evaluate and construct mine evacuation procedures and plans.

Jan 11, 1981

By Jonguk Kim

Hydrothermal sulfides were recovered from the16°50?S triple junction area in the North Fiji Basin, at a water depth of ca. 1900 m. The chimney samples can be divided into three groups according to their major metal contents: 1) type 1 (Fe-Cu-rich), 2) type 2 (Fe-Cu-rich with minor Zn), and 3) type 3 (Zn-rich) chimneys. Type 1 chimneys are mainly composed of chalcopyrite and pyrite, and are enriched in elements commonly precipitated under high temperature conditions (> 300°C), such as Cu, Co, Mo, and Se. Type 3 chimneys consist dominantly of sphalerite and marcasite with traces of pyrite and chalcopyrite, and are enriched in elements commonly associated with low temperature (150 to 250°C), such as Zn, Cd, Pb, As, and Ga. Type 2 chimneys show mineralogy similar to that of type 1 chimneys but their metal contents range between type 1 and type 3 samples. Trace element compositions of basaltic rocks indicate that magma generation in the triple junction area was influenced by two different sources: N-MORB and E-MORB. Sulfur and lead isotope patterns of the hydrothermal chimneys show distinct differences between the type 1 and type 3 chimneys. For example, type 1 sulfides are depleted in 34S and have lower Pb isotope ratios compared to type 2 and 3 chimneys. The d34S values (0.4 to 5.6?) of chimney sulfides are typical range of sulfur isotope composition of sediment free mid-ocean ridge hydrothermal system. Type 2 and 3 sulfides show slightly higher d34S values than type 1 sulfides, which can be explained by subsurface boiling of hydrothermal fluids followed by mixing of saline solutions with ambient seawater. Lead isotope compositions of the chimneys indicate that the Pb in type 1 and type 3 chimneys originates from hydrothermal leaching of N-MORB and E-MORB volcanic rocks, respectively. Considering that both these chimneys are located in the same area, our model suggests that type 1 chimneys formed by hydrothermal circulation controlled by an intrusion of N-MORB magma; the fluids were not influenced by boiling during the formation of type 1 chimneys. In contrast, type 3 chimneys are younger and were formed by hydrothermal circulation closely related to E-MORB magma intrusion at shallower depth. The necessary conditions for phase separation of the hydrothermal fluids that formed the type 3 chimneys might result from changed P-T conditions in a shallower high-temperature reaction zone beneath the hydrothermal venting site. Under these conditions the fluid may cross the two-phase boundary for seawater, ensuring subcritical phase separation. The mineralogical and geochemical properties of these chimneys suggest that type 1 and type 3 chimneys were probably black smoker and white smoker, respectively. Type 2 chimneys might be product of low-temperature replacement of type 1 chimneys, which is related to formation of type 3 chimneys.

Jan 1, 2005

By Joseph J. Yancik, Leonard L. Felts, George B. Clark

LITTLE information has been published concerning the actual or useful amount of energy obtained from explosives when they are used for blasting. To provide more data on this subject, 8-in. neet cement cubes were blasted in a steel plate box and the breakage energy evaluated by comparison with drop crusher results using Rittinger's,1,2 Kick's: and Bond's' theories. Theories by Bond and Wang," Lineau," Hatch,' Roller,' and others0-11 were not considered applicable. Those used in this investigation may be expressed mathematically as follows: Rittinger Theory E = KrS2, pet,—--------------- [1] 40 d1 d0 Kick Theory TV 1 1 E = =K, [S pct. log1/d+ 100 log d0] [2] 40 d1 L. L. FELTS, Junior Member AIME, was formerly a Graduate Assistant, University of Illinois. G. B. CLARK, Member AIME, was formerly Professor of Mining Engineering, University of Illinois, and is now Chairman, Dept. of Mining Engineering, Missouri School of Mines and Metallurgy. J. YANCIK, Junior Member AIME, is a Graduate Assistant, Missouri School of Mines and Metallurgy. This paper, TP 4213A. is revised from a thesis submitted by L. L. Felts in partial fulfillment of the M.S. degree in Mining Engineering, University of Illinois, 1954. Discussion may be sent (2 copies) to AIME before May 31, 1956. Manuscript, Dec. 27, 1954. Bond's Third Theory E = Kb [S pct. log1/d1+ 100 log d0] where E — energy used in breakage, d, = average diameters of screen fractions of broken cement particles, pct, = percent weight of diameter d;, and 2 pct, = 100 pct. The above equations were employed to determine values of Kr, Kk, Kb for particle size distributions similar to those obtained from the blasting curves. These values were in turn used as a means of approximating the specific energy of the blasted material. The efficiency of crushing by blasting was then determined by dividing this specific energy by the maximum available work of the explosive per gram of cement. Knowledge of the failure of solids under static loads is developing very rapidly. Their failure under impact loads has been investigated and reported in the literature only to a limited extent. Theories of failure' hnder static loading include 1) the maximum stress theory, 2) the maximum strain theory, 3) the theory of constant elastic strain, 4) theory of maximum shearing stress, and 5) Mohr's theory of strength. All of these have elements of usefulness in establishing an applicable theory for blasting. For highly strained and compressed material such as that surrounding a detonating charge of high explosive, the Mohr theory and its implications appear to be the

Jan 1, 1957