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By T Laurencont

The Queensland Department of Natural Resources and Mines (NR&M) accepted the environmental liability for the Mount Morgan minesite in 1991. This gold and copper mine located 38 km southwest of Rockhampton had been in operation for more than 100 years. Mining leases are still active while the feasibility of a reprocessing operation is carried out. From 1993, NR&M have operated a seepage interception pumpback system which has reduced the chronic load of acid rock drainage (ARD) to the adjacent Dee River, a tributary of the Fitzroy River system. In early-2000, it was recognised that a ten-year plan for rehabilitation of the mine, along with significant funding would be required if a long-term sustainable rehabilitation strategy was to be implemented at this minesite, ultimately making the seepage interception system redundant. A rehabilitation planning workshop coordinated by NR&M was held in May 2000. This workshop brought to the mine a range of technical expertise and key stakeholder groups. This workshop provided a technical forum for review of studies undertaken and to identify priority projects, which needed to be undertaken in the future to more clearly define contaminant sources on the mine site and to prioritise the rehabilitation program. Stakeholder consultation enabled the project team to be clear on community and other government (eg Environmental Protection Agency) expectations. It is clear that the future of the economically depressed community of Mount Morgan will be reliant on the careful management of the significant heritage values at the minesite for tourism, education and research. In the first year (2000), priority studies were initiated, which included whole of site studies on water balance, contaminant sources and a review of occupational health and safety and environment of the minesite and conservation and heritage scoping. Baseline vegetation and soil materials surveys were also initiated. The outcomes of these studies helped to shape the work program for the following year, while providing a basis for an additional funding application, which was granted. The second year of studies focused on building upon the existing baseline studies completed. In addition, studies focussed on addressing the priority issue of contingency planning for overflow from the open cut pit in accordance with best practice management of release water and river monitoring strategies were completed. Completion of these projects over the next six months will enable the optimal rehabilitation strategy to be defined in 2003. This optimal strategy may require the use of a water treatment plant for part of the ten-year program to provide space in the open cut pit for the deposition of ARD contaminant sources (eg waste dumps, tailings). It is envisaged that some sources will be rehabilitated insitu while others will be relocated to the open cut pit. Implementation of this strategy will be dependent on subsequent funding.

Jan 1, 2003

By F. Melo, M. Pawlik

Fine coal aggregation in concentrated coal-water suspensions was studied through steady-state rheological measurements. The wettability of coal particles was modified by dry oxidation of a bituminous coal. The degree of hydrophobicity was monitored using film flotation. The rheological measurements revealed that hydrophobic coal particles aggregate in aqueous suspensions, over a wide pH range, and produce slurries with high yield stresses. As the degree of coal oxidation increases, the coal particles become more hydrophilic and such suspensions exhibit lower yield stresses values. The yield stress values for suspensions prepared from oxidized/hydrophilic coal become a strong function of pH and correlate well with the surface charge as determined through electroacoustic measurements. These results indicate that the yield stress of aqueous suspensions of coal, and hence coal particles aggregation, is strongly affected by the wettability of coal particles.

Jan 1, 2004

By M I. Pownceby, W J. Bruckard

A South African chromite sample was subjected to characterisation and physical beneficiation testwork in order to assess its suitability to produce an upgraded chromite product with a silica level of <1 per cent SiO2 and a chromite level of >46 per cent Cr2O3. Chemical analysis by XRF indicated that the bulk material had a Cr2O3 content of 40.5 per cent w/w and a SiO2 content of 7.5 per cent w/w. Combined XRF analysis with sizing data showed that the majority (~70 per cent) of the siliceous material was present in the +600 ¦m fraction while ~70 per cent of the Cr2O3 was in the -600 ¦m fraction. Qualitative mineral phase analysis by XRD indicated that the sample mineralogy was dominated by ferrian-magnesiochromite spinel ([Fe,Mg]Cr2O4) and aluminium-chromite spinel (Fe[Al,Cr]2O4), with enstatite (Mg2Si2O6), haematite (Fe2O3) and anorthite (CaAl2Si2O8) as minor phases. SEM and EMP characterisation showed that the silica was mainly present in discrete enstatite grains but lesser amounts were also present as the high silica phases plagioclase and enstatite on the edge of grains. Wet gravity separation performed on sized fractions (-600+106 ¦m size range) from the -2.0 mm (primary) ore as well as the ore that had been crushed to pass 0.6 mm (secondary) showed that stage crushing of the -2 mm +600 ¦m fraction followed by wet gravity separation upgraded the Cr2O3 grade at 70 per cent Cr2O3 recovery from ~44 per cent Cr2O3 in the primary material to 46 per cent Cr2O3 (the target grade specification) in the secondary sample. It also showed that the stage crushing of this +600 ¦m fraction at a cut-off level of one per cent SiO2 increased the cumulative Cr2O3 recovery from ~50 per cent in the primary material to ~63 per cent in the secondary sample in the same size range of -600+106 ¦m. The extent to which additional crushing could lead to improved chromite recoveries at the required chromite and silica grade target specifications was not determined. This experience demonstrates that prior to processing test work being conducted, a thorough understanding of the chemistry and mineralogy of the test sample promotes efficient and targeted processing research.

Jan 1, 2008

By H. H. Johnson, Eric W. Johnson

A newly developed ac technique was used to measure the electrical-resistivity changes associated with both cyclic stressing and subsequent annealing of high-purity and OFHC copper. The early stage of push-pull fatigue was explored at 293° and 4.2°K, with stresses chosen for an expected fatigue life of 2 to 8 X 105 cycles. For high-purity copper, fatigue at 293°K was accompanied by an initial increase in resistivity. After about 5000 stress cycles, no further increase was observed. Annealing studies showed that the resisticity increment could be accounted for solely in terms of an increased dislocation density, to about 2 x 1010 disl pev sq cm. Fov 42°K fatigue, however, the resistivity increased continually, with no evidence of saturation, as approximately (No. of stress cycles)1/4. This is intevpreted to result from both dislocation multiplication and vacancy production; estimated vacancy concentrations are 10-4 to 10-9. The continuous resistivity increase was inhibited by interspersed annealing treatments at 293°K, which also raised the level of the stress-strain curve. These effects are attributed to annealing of vacancies into dislocation loops. In contrast, OFHC copper fatigued at 4.2°K showed saturation of electrical vesistivity after only a few hundred cycles, and it appeared that vacancies annealed out completely between 77° and 293°K. MaNY observations suggest that both dislocation multiplication and point-defect production are important aspects of fatigue deformation. An increase in dislocation density with number of stress cycles has been demonstrated or implied by several experimental techniques, including optical metallography with lithium fluoride,' electron microscopy with copper,2 thermal conductivity of Cu-Zn alloys,3 and peak-stress measurements during cycling at constant plastic strain amplitude.4,5 At low amplitudes and/or stresses saturation of fatigue hardening and dislocation density occurs early in the expected fatigue life,4-6 frequently by a few hundred cycles. The evidence for point-defect production is somewhat less extensive, but it is nonetheless persuasive. Rosenberg7 reported briefly that, for equal stresses, copper fatigued at liquid-hydrogen temperature displays an increase in electrical resistance some ten times that which accompanies unidirectional stressing. Further, the extra resistance saturates after a few hundred stress cycles, and does not anneal significantly until a temperature of about 170°K is attained.' Point-defect production during fatigue is also suggested by 1) the softening during cyclic stressing of initially strain-hardened polycrystal-line copper9, 2) the strong temperature dependence of the yield strength of fatigue-hardened mono- and polycrystalline copper,579"0 and 3) the influence of intermediate annealing upon slip-band formation in copper fatigued at 90°K.8 It is evident that characterization of the fatigued state requires a knowledge of both point-defect and dislocation densities. For this purpose electrical-resistivity measurements are useful, but they are difficult to perform because of geometric limitations. Push-pull fatigue is necessary to obtain a macroscopic ally uniform deformation over the specimen cross section; however, the specimen must then have a low slenderness ratio to avoid plastic buckling, while the standard dc resistivity technique requires a very slender specimen for

Jan 1, 1965

By H. H. Kellogg, T. R. Ingraham

Three anhydrous zinc sulfates have been identified. They are: ZnSO,(a), stable below 1007°K; ZnS04(/3), stable above 1007OK; and ZnO.ZZnSO,. The decomposition pressure of each sulfate has been measured and the following relations calculated from the results: The measured data have been combined with the known themzodynamic properties of ZnS, ZnO, Zn, SO,, and SO, to yield univariant and bivariant phase diagrams for the system Zn-S-0. The application of these diagrams to problems of roasting zinc concentrates is discussed. PRESENT knowledge of the equilibrium chemistry of the system Zn-S:O, so essential to scientific control of zinc-blende roasting, is deficient in several respects. A basic sulfate of zinc has been reported by many investigators, but its composition is in doubt and its thermodynamic properties have not been measured. Our scanty knowledge of the properties of the normal sulfate at roasting temperatures is based on old and inexact studies. A crystal transformation in the normal sulfate, involving a sizable heat effect, has been noted by several investigators but has never been quantified. The studies reported herein were designed to fill these gaps in knowledge and to make possible a reasonably complete thermodynamic description of the system Zn-S-0 at roasting temperatures. The results make possible a far more exact understanding of zinc roasting chemistry than was hitherto possible. EXPERIMENTAL The principal experimental technique employed was the measurement of the total gas pressure developed from decomposition of either ZnSO, or ZnO2ZnS0, in a closed system. The apparatus employed was essentially the same as that described by Warner and lngrahaml and Flengas and tngraham.' The unusual feature is a flexible Pyrex bellows used to separate the reaction gas mixture (S03, SOz, and 02) from the mercury in the manometer and thus prevent corrosion of the mercury by SO3. The bellows operates without evidence of hysteresis and, barring accidents, may be used for several months. The response of the bellows varies from manometer to manometer, but is generally of the order of 0.6-cm displacement for a l-cm change in pressure. The response for all manometers has been linear over the range 0 to 1 atm. To prevent condensation and polymerization of sulfur trioxide, the bellows-manometer was maintained at 100.0" * 0.5" by a circulating bath of silicone oil. The connecting tubing between the manometer and the furnace was maintained at 100.0" * 0.2"* also, with a stirred air bath. The open end of the manometer was connected with a chamber which was maintained at a known pressure. The pressure could be varied so that readings of the height of the mercury column, measured with a cathetometer, could be made with both rising and falling mercury column. By using the average of the readings, any error due to sticking of the mercury column was minimized. The sample was placed in a 2-in.-long platinum boat formed into the shape of a half-cylinder. The sample was prepared as a fine powder and pressed against the inner wall of the platinum boat. A silica thermocouple well, covered with a platinum sleeve, was placed lengthwise along the center of the boat. The temperature of the sample was read on a calibrated Pt-10 pct Rh theirnocouple. The temperature profile of the furnace was constant within ± 0.2 C over the length of furnace occupied by the sample. The horizontal furnace used to contain the sample and the quartz reaction tube was electrically heated, with the control thermocouple placed 1/4 in. from the windings. A Brown Pyro-O-Vane temperature controller with a 3-sec response time was used to maintain the furnace temperature. Before beginning a run, the apparatus, including the bellows manometer, was thoroughly washed in sib with acid and distilled water to remove any

Jan 1, 1963

By W. W. Adams

"Accidents at coal mines in the United States in September 1926 resulted in the loss of 170 lives, according to information furnished to the Bureau of Mines, Department of Commerce, by State mine inspectors. Of this number 125 occurred at bituminous mines and 45 in the anthracite field. As the bituminous output of coal for September was 48,976,000 tons, and the production of anthracite was 8,444,000 tons, the fatalities reported represented a death rate of 2.55 per million tons for bituminous mines and 5.33 for anthracite mines. The industry as a whole showed a fatality rate of 2.6 per million tons of coal produced in September 1925 the rate for bituminous coal mines was 3.29; the anthracite mines were idle in September last year due to a strike.During the first nine months of the present year 1,759 men have been killed by accidents at the coal mines throughout the United States, resulting in a death rate, per million tons of coal produced, of 3.76 as compared with 3.95 for the same period in 1925. Reports for bituminous mines alone showed a nine-month fatality rate of 3.53 which was about the same as the corresponding rate from January to September in 1925. The anthracite rate for the present year, however, shows a decrease of about 17 per cent; that is, the rate has declined from 6.36 to 5.28.Records covering the first nine months of 1926 show that 12 ""major"" disasters, that is, disasters in which five or more men were killed at one time, have resulted in the loss of 271 lives as against 10 disasters and 199 deaths in the corresponding months in 1925. The fatality rates for these disaster: alone were 0.579 and 0.465, respectively. One major explosion occurred at Tahona, Oklahoma, on September 3, 1926, in which 16 men lost their lives.An analysis of the principal causes of the 1,759 fatalities for the first nine months of the present year shows a reduction in the death rate as a whole and for falls of roof and coal and for explosives, but an increase in haulage, gas and dust explosions, and electricity. The comparative rates for these accidents were:"

Oct 1, 1926

By Albert L. Toenges

The fuel requirements for industrial and domestic consumers in New England, in normal times, are supplied almost entirely by fuel oil, anthracite, and bituminous coal. A shortage of these fuels occurred during the war. This shortage was caused by: (1) The diversion of oil to military uses; (2) the use of railroad and boat facilities that ordinarily would transport coal for the transportation of war material; (3) the inability of coal mines to apply the increased demand for coal which resulted from the conversion from oil to coal by domestic and industrial consumers; and (4) the increased demand of war industries for coal. Because of this Critical fuel situation, an investigation was made of possible sources of fuel nearer to points of consumption in New England.

Jan 1, 1948

By M. A. Reyes, J. Bickson, J. R. Srednicki

Federal laws mandate the presence of refuge alternatives (RAs) in underground coal mines and provide guidelines for harmful gas removal. After an explosion, harmful gases can follow survivors into an RA and can reach a deadly concentration. Investigators from the National Institute for Occupational Safety and Health (NIOSH) tested 12 ventilation system configurations (VSCs) to purge sulfur hexafluoride (SF6) tracer gas from a built-in-place (BIP) RA in the Experimental Mine in Bruceton, PA. Using a borehole air supply, tests were conducted with airflows of 750 and 1,000 standard cubic feet per minute (SCFM). Excluding a VSC where the air inlet was 1.98 m (78 in) from the air outlet, the results show that the time to purge the SF6 concentration from 1,000 parts per million (ppm) to 25 ppm ranged from 26 to 29 minutes for an airflow of 750 SCFM and ranged from 18 to 23 minutes for an airflow 1,000 SCFM. Mines can use this information to reduce purge time in case of a mine disaster and determine which VSC to avoid. INTRODUCTION After a fire or an explosion in an underground mine, the concentration of carbon monoxide can reach a fatal level. In a mine disaster, the main goal is to escape. However, if miners cannot escape, they have to migrate to a location to find refuge and, if possible, communication, and a borehole source for a fresh air supply. To ensure survival of the miners, Federal laws mandate the presence of a refuge alternative (RA) within 304.5 m (1,000 ft) of the nearest mining face. In addition, Code of Federal Regulations Title 30 CFR Section 7 sets the guidelines for personnel tracking, RAs, and the removal of noxious gases in underground coal mines (Mineral Resources, 30 CFR, 2014). Federal regulations define an RA as a protected, secure space with an isolated atmosphere and integrated components (Mineral Resources, 30 CFR, 2014). There are three types of RAs: portable metal RAs, portable tent-type RAs, and built-in-place (BIP) RAs. A BIP RA is constructed using masonry or other stoppings to create an isolated space in a mined-out area. Blast doors are installed in one or both of the stoppings for the miners to enter the RA. Furthermore, RAs must provide the following components: breathable air, food, water, air monitoring, and harmful gas removal. Federal regulations have emphasized the mandate for breathable air and harmful gas removal in the following words: “Purging or other effective procedures shall be provided for the airlock to dilute the carbon monoxide concentration to 25 ppm” (Mineral Resources, 30 CFR, 2014). The purpose of the mine ventilation system is to provide a sufficient quantity of air to the underground mine workings, to dilute methane and other contaminants, to maintain a suitable working environment, and to prevent accidents from happening (Cheng and Wei, 2014). In addition, the organic nature of coal amplifies risks from poisonous gases, explosions, and fires (Engström et al., 2018). If an underground coal mine ventilation system fails after an explosion, the concentration of toxic gases, most notably carbon monoxide, will increase. Carbon monoxide is dangerous because it creates tissue hypoxia by binding to hemoglobin (Özmen and Aksoy, 2015).

Jan 1, 2019

By J. L. Fuller

Mine ventilation surveys were conducted by Technical Support engineers during June 1987 and September 1989 at the Dutch Creek Hine. The mine is located near Redstone, Colorado and is operated by Mid-Continent Resources, Inc. Following completion of the initial survey, several recommended improvements were incorporated into the ventilation system by mine personnel. The second survey permitted documentation of the effects of those changes on the ventilation system performance. This paper describes the processes involved in the implementation of the recommended changes. The substantial improvements in system efficiency and reduced operating costs are also described.

Jan 1, 1991

By D. C. Torre

Ever since the early '60's there have been efforts to develop and produce reliable continuous haulage systems capable of taking the full output of a continuous miner and transporting it, in one continuous flow, from the face to the tipple. The majority of the continuous haulage systems which have been most successful to date were placed in operation in the early '70's and were all of the bridge conveyor type, usually consisting of bridges and mobile bridge carriers attached at the inby end to the continuous miner and at the outby end to the panel conveyor itself. These successful bridge conveyor type continuous haulage systems were, and continue to be, either of the belted or chain type. Undoubtedly, the most successful belt type continuous haulage systems are those manufactured and marketed by the Long-Airdox Company. Until early 1977, the most successful chain conveyor type haulage systems were manufactured and marketed by Jeffrey Mining Machinery Division. In 1977, Long-Airdox reentered the market with a chain type system to complement its belted system. I say "reentered" because the original bridge type continuous haulage system was introduced to the coal mining industry in the early '60's by the old Long Company, a predecessor of the current Long-Airdox Company, and those original units were of the chain conveyor type.

Jan 1, 1981

By Peter Cunningham

Teck Facts and Figures ? Employ more than 14,000 around the world, includes: ? Over 8,600 in Canada (close to 7,800 in British Columbia) ? Close to 2,000 in Chile ? Over 2,700 in Peru ? Approximately 600 in U.S. ? Hundreds more in exploration around the world ? Total capital expenditures ? $1.8 billion for 2012

Aug 1, 2013

By Patricia J. Dillon

Established in 1997, Prospectors and Developers Association of Canada Mining Matters (PDACMM) is a registered charitable organization dedicated to providing Earth Science education to grade school students. Through the development and distribution of educational programs and materials, PDACMM informs students of Canada?s geology and mineral resources as well as promoting awareness of the importance of rocks, minerals, metals and the mining industry. PDACMM also co-ordinates with other charitable organizations and government agencies to provide support or to work in tandem with them and reach as many young people as possible. Although PDACMM became a charitable organization in 1997, it actually has its? roots in 1994 when the Grade 7 program Mining Matters/Une Mine de Renseignments was developed to help teachers bring geology and mining into the classroom. Over 2100 of these resource kits were distributed to teachers across Ontario during the period 1995 to 1998. This unit won national recognition when it received the prestigious Partnership Focus award from the Conference Board of Canada for ?promoting science literacy for the world of work?. Due to the phenomenal success of these kits, a Grade 4 kit entitled Deeper and Deeper/Aller au Fond Des Choses was developed in 1998 in response to the opportunity provided by the new Ontario curriculum released at that time. Over 2600 of these kits were distributed throughout Ontario and there is a substantial waiting list of teachers hoping PDACMM will be able to raise enough money for another print run.

May 1, 2002

By R. A. Busch

Sludge ponds of fine coal, fine soil, and water created behind large piles of coarse coal waste can present hazardous conditions; this is a Bureau of Mines report of the subsurface physical property data developed on the fine coal waste materials that make up the core areas of two typical coal waste impoundments. These data will help define more completely the engineering properties of this problem area of the structures. Additional calorific, chemical, and mineralogical data were obtained on the coal waste materials to determine possible uses or development of byproducts. The possibility of extracting alumina from coal ash was also investigated.

Jan 1, 1975

By Donald A. Dahlstrom

In 1961 it became increasingly more evident that significant technological changes must be made in minerals beneficiation. The status quo in technology in today's competitive markets will not be tolerated. Higher productivity, greater recoveries and newer and better products must be sought and incorporated with lower operating costs. In other words, "when you build a new mill, don't build an old mill." Recommended reading for proof of the vital necessity of new methods and techniques is in the September 1961 issue of MINING ENGINEERING. The article by one of MBD's cherished and true veterans, E. H. Rose, is entitled "Iron Ore: The BIG Picture". Because of new beneficiation methods for iron ores, nearly 47 million tons of plus 60% iron concentrates will be produced and another 40.1 million tons of capacity will be under construction or committed. Thus in a relatively short period, 87 million tons of this high-grade iron concentrate with record low silica will be available. As Mr. Rose also points out, "in gross tons, without allowance for better grade, this is the equivalent of almost exactly two-thirds of all iron ore used for all purposes in the U.S. and Canada in the peak years of 1953 or 1957." Needless to say, improved mineral beneficiation methods have had far more reaching effects than just the production of higher grade iron concentrate.

Jan 2, 1962

By W J. Dyck

Past attempts to establish an integrated steel industry on the West Coast of Canada have been unsuccessful for at least three reasons: (a) insufficient reserves of a suit-able iron ore supply; (b) a relatively small local market for iron and steel; and (c) the non-availability of an economical process to produce primary iron on a small scale. Today, the reserves of economically mineable ore deposits are ample to support a steel production of 100,000 to 200,-000 tons per annum. The local market has increased so as to be able to absorb most of this capacity. With-in the next few years, a small integrated steel plant, including the "H-Iron" direct iron ore reduction process for electric furnace charge stock production, will be able to show a profitable operation. Introduction

Jan 1, 1963

By L. Adler

Any design of a mine roof in bedded deposits which ignores differential deflections at the supports can quickly lead to dangerous overstressing. As illustrated by the typical case presented on page 1029, if one support deflects only 0.1 in. more than its neighbor, stresses 50 pct higher than anticipated will result. Previous studies1-3 have assumed that the mine roof acts like a beam clamped at both ends, with no deflection or slope' at the supports (Fig. 1). Essentially this assumes:5* ?A = ?, = ?b and dA = dB = 0. However, a host of conditions can cause differential sag or deflection at one support, and even an apparently trivial deviation can have major consequences in the beam: variations in spacing, size, and material of supports; local settlement due to floor heaving or lower mine openings; non-uniformity in wedging or bolt tensioning.8 In one particular instance concrete pillars, poured in place, shrank away from the roof upon setting.' When understood and controlled, differential sag becomes a definite advantage, permitting the transfer of loads from one support to another. Where one support is a bolt or prop and the other a pillar or rib, the load on the former can be reduced while the load on the pillar or rib is increased. This means two things: 1) savings can be effected by using smaller artificial supports such as props and bolts and 2) loads can be increased in the rib to induce eventual failure for longwall caving. In this last respect, another factor underlying coal bumps becomes understandable. The design of a mine roof by beam theory assumes that when a bedded formation is homogeneous, iso-tropic, and elastic, it will act like a uniformly loaded beam, fixed or clamped at both ends,6 s in Fig. 1.'" These are important assumptions. The uniform load is the weight of the beam itself or its body load. For a beam of unit width, this load becomes per foot: w = yhb. w = uniform load y = specific weight, lb per ft3 h = depth or thickness of beam, ft b = width of beam, ft Since b = 1, w - yh [ 1 ] A beam fixed at both supports is statically indeterminate to the second degree, that is, two equations in addition to those for static equilibrium are required to determine the reactions at the supports. These equations result from the geometry. From statistics: ^F,= 0;RA + RB — wL = 0 [2] wU S,M0 = 0 ; MB H------------MA — RBL = 0 [3] it Using superposition in which the left support is kept intact and the right one resolved into an active force and moment, we obtain from Fig. 2:" ?b = ?m + ? b2 + ?B3 = 0 dB = dB1 + dB2 + dB3 / wU Ma RBL \ V d°=(-r+-2—r)-zr [5] Where: Subscripts refer to position on axis of beam R = force, lb M = moment, ft-lb w = uniformly distributed load, lb per ft L = span, ft 0 = slope, radians d = deflection, ft E = Young's Modulus, lb per ft" 1 = Moment of inertia about the neutral axis, ft4 Solving Eqs. 2 through 5 for the four unknown reactions, we obtain: 2 L3 Fig. 3 is obtained by plotting MA and M, vs d, and then plotting RA and R, vs d, and adjusting their scales to coincide with MA and M, vs d,. It should be noted in Eqs. 6 and 7 that as dB increases, MA increases the same amount that Mn decreases. Eqs. 8 and 9 show a similar relationship between RA and R,. Proof that MA is the maximum absolute moment occurring in the beam is as follows: where M, = moment at x wx-When where M,., = maximum positive moment Eqs. 6 and 8 into above: For small dn: MA> M,,,

Jan 1, 1960

By R. W. Wittkopp

The Relief Canyon gold deposit is located approximately twenty- four kilometers east of the town of Lovelock in Northwestern Nevada. At Relief Canyon, the Late Triassic Grass Valley Formation has been thrust over the Late Triassic Natchez Pass Formation. The lower plate Natchez Pass Formation is composed of massive gray dolomitic limestone which is locally silty and sandy. The upper plate Grass Valley Formation is composed of shale, siltstone, argillite and quartzite. Gold mineralization at Relief Canyon occurs along the highly brecciated thrust contact between these two formations. Weak gold mineralization often occurs above the thrust in the Grass Valley Formation, however, most of the ore grade mineralization is present below the thrust in the brecciated Natchez Pass Formation or where the two formations are mixed as a result of movement along the thrust. Several types of gold mineralization have been recognized at Relief Canyon. A large amount of the ore is present as a jasperoid breccia, however, significant amounts of ore occur as a mixed breccia composed of upper plate rock, fresh limestone, clay and jasperoid breccia. Gold ore also occurs in a fresh limestone breccia which contains a large amount of clay. Below masses of jasperoid, jasperoid stringers containing gold extend into the fresh limestone. Gold mineralization at Relief Canyon appears to be truly disseminated and has the typically associated trace elements of arsenic, antimony, mercury and fluorine. The Relief Canyon deposit can be classified as a structurally controlled disseminated gold deposit. Other deposits that fit this classification are the Pinson and Dee gold mines.

Jan 1, 1984

By Rick Valenta

Is the Labrador Central Mineral Belt part of the IOCGU Club? ?I sent the club a wire stating, PLEASE ACCEPT MY RESIGNATION. I DON'T WANT TO BELONG TO ANY CLUB THAT WILL ACCEPT ME AS A MEMBER.? GrouchoMarx Acknowledgements ?Ian Dunlop, Matt Lennox-King, Mark O?Dea, Dave Barbour, Derek Wilton, Lucas Marshall, Roland Butler, Mark King, Sally Howson,

May 1, 2006

By Alan Morris

Brass rod for use in automatic screw machines is one of the major products of the brass mills. A large tonnage is consumed each year in the manufacture of an endless variety of finished articles and parts for assembly into other products. The machines that do this work have been developed to a high degree of efficiency, and their manufacturers are still striving for higher production speeds. This development has brought with it a demand for brass rod to meet the increasingly difficult machining conditions. This paper is a report of the results of tests undertaken in an effort to obtain a measure of the effects of such variables as lead content, microstructure and cold work on the machinability of brass rod. It is realized that the term "machinability" has never been well defined. Power, finish and tool life all should have a place in such a definition. In practice it is sometimes found necessary to sacrifice some of these in order to obtain a rod that will not tend to throw up a burr, or will give a burr which is easily tumbled off. Again a specially stiff rod may be necessary because of long overhang from the chuck or a rod of special characteristics is needed to take a given broaching operation without chattering. It is only through an understanding of the effect of basic mill operations on the machinability of the rod that such problems as these may be solved most economically. It is hoped that this paper, together with the discussion it may evoke, will prove to be a basis for such an understanding. Testing Apparatus The testing apparatus used was of the single-tooth milling cutter type, in all respects sirnilar to that first used and described by Airey and Oxford1 and again by Boston. The machine used in the present test is shown in Fig. 1. It is an ordinary milling machine in which a heavy pendulum has been substituted for the pulley on the arbor. If the pendulum is raised

Jan 1, 1932

By Charles E. Green

This study was conducted to identify, evaluate, and develop procedures to minimize or eliminate safety hazards associated with the electrokinetic dewatering process, such as toxic and volatile gas emissions, electrical shock, accidental detonation of explosives, electrochemical corrosion of metal appurtenances, and possibly others. The identification, evaluation, and procedure development were based on pertinent literature, existing safety regulations, interviews with mine personnel, physical and mechanical properties of the electrokinetic dewatering process, and physical characteristics of the material to be dewatered. The electrokinetic dewatering process was described in detail. Potential safety hazards were identified and evaluated. Procedures and/or safety regulations were developed to minimize or eliminate the potential hazards. It was recommended that the electrokinetic dewatering process when used in deep metal mines be properly monitored to obtain additional information on other potential hazards that possibly exist.

Jan 1, 1979