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By David Gatto, Nicolas A. Willig-Friedrich, Jeff A. Segar

"With the increasing demand for electricity, and requirements to generate this power cleanly, more producers are turning towards hydroelectric power. A project in Iowa currently under construction will retrofit an existing Army Corps of Engineers Dam with a hydroelectric generating facility. The design of the intake structure requires the construction of a diaphragm retaining wall. The T-shaped diaphragm wall elements retain an unbalanced fill height up to 21.3 meters (70 feet). The design required a single production element be constructed, tested, and instrumented prior to beginning construction of the rest of the elements. The testing included extensive concrete mix design development, coring, and Cross Hole Sonic Logging. Installed in the other production elements are 10 inclinometers, 4 pressure cells, 4 piezometers, and multiple survey monuments. Some of these instruments require manual readings while most are monitored by an automatic data acquisition system and instantly transferred to a website providing real-time graphical display of the data. This paper will discuss the construction of the diaphragm wall from the contractor’s perspective when working on a site only large enough for two service cranes and a Hydromill excavation unit, without any storage area on the construction pad, describe the testing performed and the quality control procedures employed, and investigate the information gained from the automated instruments providing real time measurement readings.INTRODUCTIONIn 1964 the US Army Corps of Engineers (USACE) constructed a dam by raising the surface grade approximately 34.7 meters (114 feet) above the natural ground line using a combination of fill types. This earthen dam stretched slightly longer than a mile intersecting the flow of the existing Des Moines River. The center of the dam is a concrete spillway consisting of 12 monolith sections. On each end, three of the sections extend 42.6 meters (140 feet) from the spillway to transition from the concrete dam sections to the earthen dam section. Adjacent to these three sections on the East end is where the contractor is actively working, for a private developer under a lease agreement with USACE, on this site to convert it to a Hydroelectric facility licensed for an output of 36.4 megawatts, enough power to meet the needs of about 18,000 homes. To create this, a powerhouse structure, housing two vertical Kaplan turbines is being added to the downstream side of the existing dam. From the power house, two penstock pipes will be cored through the existing concrete monolith sections and connect to the Intake structure on the upstream side of the dam. Protecting and reinforcing the dam adjacent to the intake structure is a new diaphragm wall. This wall is constructed using T-shaped diaphragm elements, acting as a cantilever retaining wall for the intake platform. Figure 1 shows a diagram of the upstream features of the construction."

Jan 1, 2016

By Liu Xiao, Lin Xiaoying, Ma Geng, Tao Yunqi

"The soft coal in China has several issues, low permeability, high gas content and prone to gas outburst, making it difficult to drill and improve its permeability. Using the rock mass loading test, the relationship of stress-strain-permeability-time of sandstone was investigated. The fracture network that communicates with coal seams to provide the pathways for gas migration can be established by drilling, hydraulic fracturing or loose blasting in the roof and floor about 5m from coal seams top and bottom. The gas migration can be summed up as ""desorption-diffusion-two levels of seepage"" for gas drainage of the quasi-protection-layer in hard coal seams, and ""desorption-two levels of diffusion-seepage"" in soft coal seams. In the quasi-protection-layer, different gas drainage techniques were used during different stages of mining. Gas drainage of the quasi-protection-layers was demonstrated at Hebi Zhongtai Mining Co., Ltd. The results showed that the technique promotes the gas seepage flow, and improves the efficiency of drainage. The influence zone of the quasi-protection-layer was 1601 m2 and the average amount of gas drainage was 328m3/d, with the maximum being 661m3/d. INTRODUCTIONGas is one of the main factors that cause disasters and influence safety production in underground coal mines (SACMS, 2009). The coal-bearing formations in China have been subjected to multiphase tectonic stress field during the period of geological evolution. Coal body was often crushed into fragmented coals and mylonitic coals under the tectonic stresses (Su and Lin, 2009). The gas permeability of coal is generally poor. So in underground, drilling and hydro-fracturing technology have been used to improve the permeability of coal, such as intensive drilling, hydraulic punching, hydraulic cutting, hydraulic extrusion and deep hole blasting. But the gas drainage rate was low, only 23% (Liu, Li, and Li, 2006; Li and Wei, 2006; Henan Polytechnic University, 2008; Ma, et al., 2014; Lin and Wu, 2009; Zhou, et al., 2011) that was much lower than the rates achieved in the main coal-producing countries such as America and Australia whose average rate of gas drainage was 50%. The protection layer mining technology has been used widely as an effective technique to improve the permeability of coal in the mining of multiple coal seams (Cheng and Yu, 2003; Diaz and Gonzalez, 2007; Karacan, et al., 2007; Palchik, 2003; Latham, et al., 2013; Yuan, et al., 2013; Shi and Liu, 2007; Wang, et al., 2008, Guo, et al., 2015; Lu, et al., 2015; Slazak, Obracaj, and Swolkien, 2014 ). In the single low permeability coal seam, there is no protection layer. This kind of coal seam has low permeability, high gas content, soft coal body, and is prone to serious gas outbursts. During drilling, the holes very often spray, collapse, stuck and wedged the drilling tool. The hole length drilled is difficult to reach the required minimum, the service life of holes for gas drainage is short, the construction cost is high, and the holes are difficult to maintain. The gas drainage efficiency of single low permeability coal seam, especially soft coal seams, cannot be improved even though hydro-fracture technology has been implemented, because its permeability is not improved."

Jan 1, 2016

By N. P. Goss

ASTERISM appearing in X-ray Laue dia-grams is an extremely sensitive index of changes in the internal structures of indi-vidual grains of polycrystalline metals. It indicates the existence of various forms of subboundary structure or subdivision within the grain. The present viewpoint is that the grain is made up of lattice blocks having substantially the same size and shape.1-4 When these blocks all assume exactly the same orientation, the crystal is perfect and asterism is absent. Cleaves and Thompson5 observed that the hardness of iron is virtually constant over a range of grain sizes. It is quite possible that the subboundary structure caused this, and that the latter is just as important as grain size itself. While the experiments reported here are not con-clusive, they do support this viewpoint. Metals can, under certain conditions, be prepared entirely free from asterism, but this preparation does not necessarily require long heating or slow cooling, as will be shown here. In this paper a distinction is made be-tween mosaic and distorted grain struc-tures but the difference is only a matter of degree. In a previous paper9 it was shown that the grains can become distorted even during heat-treatment. The block structure exists inherently, even in a perfect crystal; the size and shape of these crystallites, or lattice blocks, varies, depending upon a number of factors. Goetz, for example,7 has shown that a metal before crystallizing from the molten state must go through a paracrystalline phase with oriented groups of atoms about 10-14 to 10-15 c.c., or 10-5 cm. in size.

Jan 1, 1940

By N. P. Goss

ASTERISM appearing in X-ray Laue dia-grams is an extremely sensitive index of changes in the internal structures of indi-vidual grains of polycrystalline metals. It indicates the existence of various forms of subboundary structure or subdivision within the grain. The present viewpoint is that the grain is made up of lattice blocks having substantially the same size and shape.1-4 When these blocks all assume exactly the same orientation, the crystal is perfect and asterism is absent. Cleaves and Thompson5 observed that the hardness of iron is virtually constant over a range of grain sizes. It is quite possible that the subboundary structure caused this, and that the latter is just as important as grain size itself. While the experiments reported here are not con-clusive, they do support this viewpoint. Metals can, under certain conditions, be prepared entirely free from asterism, but this preparation does not necessarily require long heating or slow cooling, as will be shown here. In this paper a distinction is made be-tween mosaic and distorted grain struc-tures but the difference is only a matter of degree. In a previous paper9 it was shown that the grains can become distorted even during heat-treatment. The block structure exists inherently, even in a perfect crystal; the size and shape of these crystallites, or lattice blocks, varies, depending upon a number of factors. Goetz, for example,7 has shown that a metal before crystallizing from the molten state must go through a paracrystalline phase with oriented groups of atoms about 10-14 to 10-15 c.c., or 10 5 cm. in size.

Jan 1, 1940

By C. B. Alcock, R. G. Hudson

Sulfur pressure measurements above FeS in equilibrium with iron have been carried out by the Knudsen orifice method. A comparison is made of the weight loss of the cell per unit time obtained in the above experiments, that calculated from Richardson and Alcock's measurements of the S2 pressure above FeS in equilibrium with iron, and the dissociation constant for diatomic sulfur from spectroscopic and thermal data. EXTENSIVE equilibrium measurements have been made to determine the sulfur potentials of a great many systems of metallurgical interest.' In order to confirm their reliability, to supply data which may be of use in further thermodynamic calculations, and to establish the reliability of the Knudsen orifice in obtaining accurate thermodynamic values at high temperatures, the present investigation on the pressure of the gaseous species above FeS in equilibrium with iron has been carried out. The Knudsen orifice method has been used for these pressure measurements. Experimental The apparatus and general technique for this investigation are the same as that reported for the determination of the vapor pressure of silver.' To reduce the residual gas pressure in the evacuated chamber, a titanium getter, maintained at run temperature, was introducbed. Several types of Knudsen cells were used—platinum, zircon, and fused silica. The orifice in the zircon cell was fashioned with a small diamond wheel. The same precautions were used in this investigation as were employed in the work on Mo2S2.3 The FeS-Fe mixture was prepared by heating a sample of pure FeS,, supplied by K. K. Kelley of the Bureau of Mines, with excess iron powder, cp grade, obtained from the Fisher Scientific Co. These two substances were heated in a closed Armco iron crucible at 960°C, held in a vacuum of less than 10" mm Hg for 24 hr. The mass was then removed from the cell and ground in an agate mortar. In this form it was charged to the various Knudsen cells. Heating the Knudsen cell for an additional two days was sufficient to obtain constant rates of weight loss of the cell. The data obtained in this investigation are recorded in Table I. To establish whether or not a sulfide of iron is volatile at these temperatures, 0.7646 g of the FeS-Fe mixture, described above and containing 69.2 pet Fe, was placed in the zircon Knudsen cell. Run Nos. 103 to 109 were performed on it. After the completion of these runs, the total amount of iron left in the cell was obtained by chemical analysis. It contained 0.537 g Fe, compared to 0.529 g in the original charge. Thus, there was an apparent gain in iron of 1.5 pet. This is attributed to unknown experimental error, and the conclusion is that no sulfide of iron is volatile. The vapor pressure of pure iron4 is about one thousandth that of the sulfur pressure' at these temperatures. The vapor pressure of iron inside the Knudsen cell will be almost exactly that of pure iron, since in the iron-rich phase the percentage of sulfur is of the order of 0.025 pet.At this low percentage of sulfur Henry's law is probably obeyed by the sulfur. Thus, the iron will obey Raoult's law. In any event, the vapor pressure of iron will be lower than that of pure iron and, since the vapor pressure of pure iron is negligible compared to the pressure of sulfur, the weight loss of the cell due to the vaporization of iron should be negligible. Since neither iron nor FeS is volatile to an appreciable extent, the total weight loss of the cell must be due to the vaporization of sulfur. Discussion and Calculations On the assumption that S, is the only gaseous species present in appreciable quantity, its pressure

Jan 1, 1957

By P. T. Bissert, T. J. Lutz, J. A. Yonkey, G. T. Homce

"Underground refuge alternatives require an air source to supply breathable air to the occupants. This requires pressure relief valves to prevent unsafe pressures from building up within the refuge alternative. The U.S. Mine Safety and Health Administration (MSHA) mandates that pressure relief valves prevent pressure from exceeding 1.25 kPa (0.18 psi), or as specified by the manufacturer, above mine atmospheric pressure when a fan or compressor is used for the air supply. The U.S. National Institute for Occupational Safety and Health (NIOSH) tested a variety of pressure relief valves using an instrumented test fixture consisting of data acquisition equipment, a centrifugal blower, ductwork and various sensors to determine if the subject pressure relief valves meet the MSHA requirement. Relief pressures and flow characteristics, including opening pressure and flow rate, were measured for five different pressure relief valves under a variety of conditions. The subject pressure relief valves included two off-theshelf modified check valves, two check valves used in MSHA-approved built-in-place refuge alternatives, and a commercially available valve that was designed for a steel refuge alternative and is currently being used in some built-in-place refuge alternatives. The test results showed relief pressures ranging from 0.20 to 1.53 kPa (0.03 to 0.22 psi) and flow rates up to 19.3 m3/ min (683 scfm). As tested, some of the pressure relief valves did not meet the 1.25 kPa (0.18 psi) relief specification.IntroductionThe Mine Improvement and New Emergency Response Act of 2006 (MINER Act) was enacted in the United States in the wake of three mine explosions/fires that claimed 19 lives that year. Intended to help improve underground coal mine accident preparedness, the MINER Act includes provisions that target mine safety issues in areas such as emergency response planning, adoption of new technology, training and education, and mine safety standards enforcement (U.S. Mine Safety and Health Administration, MSHA, 2006). Section 13 of the MINER Act specifically directed the U.S. National Institute for Occupational Safety and Health (NIOSH) to provide for research into the effectiveness and viability of refuge alternatives for underground coal mines. This mandate culminated in the 2009 adoption of changes to Title 30 of the Code of Federal Regulations (30 CFR) mining health and safety regulations, requiring underground coal mines to provide mine emergency refuge alternatives and associated components, to provide a life-sustaining environment for persons trapped underground. Such refuge alternatives can be either self-contained mobile units or built-inplace facilities. The regulatory changes also include provisions establishing requirements for MSHA approval of refuge alternatives and their components, and among these provisions are numerous criteria for providing a safe breathable atmosphere under positive pressure within the refuge alternatives. One specific criterion for maintaining a safe refuge alternative atmosphere requires the inclusion of an air pressure relief valve that will activate at a maximum of 1.25 kPa (0.18 psi), or at a pressure above mine atmospheric pressure, in the refuge alternative, as specified by the manufacturer when the breathable air is supplied by a fan or compressor (MSHA, 2008)."

Jan 3, 2018

By AIME AIME

THE papers presented at the geophysics session" on Feb. 17 were concerned largely with three aspects of the science. The first ones dealt with the transmission of elastic waves through the earth, then electrical methods came in for their share of attention and finally, the question of the teaching of geo- physics came up for the first time at these meetings. Also, for the first time, there was a symposium of mining men on the practical value of the new science. The paper (T. P. 267) entitled "Seismic Propagation Paths" was presented by Donald C. Barton in the absence of the authors. It dealt with an investigation of the speed of propagation of seismic waves. The latter travel in a straight line when proceeding at a constant velocity, but if the velocity increases at increasing depths, the line of travel becomes a curve. Under the assumption of the authors this curve is found to be the arc of a circle, and their paper gives calculations for reflected and refracted wave paths under these circumstances.

Jan 1, 1930

By Raymond Henn

Yes we have also used some however, again because of the nature of the maintenance window where we have that two hour (at best) window every night and due to the numerous other crews that are wanting to get in to do the other types of maintenance activities, we don't know if we are going to be getting into that same area from one night to the next. So even with the dormancy additives that one can add to cementitious grouts (that are good up to 72 hours) we don't know if we are going to get back in to the same locations from one week to the next. That's why I use the term "the forgiving nature of the acrylamide". It allows us to ream out the hole allowing us to use the exact same hole without having to re-drill a new hole. The program started in 1997 and we have a budget between labor/materials and equipment of 4.5 million dollars on a year to year basis. For the next few years that is going to come down drastically because the program is going to be more of an ongoing maintenance program. The program started in 1997 and we have a budget between labor/materials and equipment of 4.5 million dollars on a year to year basis. For the next few years that is going to come down drastically because the program is going to be more of an ongoing maintenance program. Up to the end of 2001 we were just over 1.7 million liters of between AV100 and AV118 grout. Of that, approximately just under 200,000 was AV100 and the remaining balance was AV118.

Feb 10, 2003

By B. Orberger, J. M. Milazzo, S. B. Blancher, J. Lützenkirchen, J. L. Dubos

"Industrial dusts are commonly agglomerated using empirical recipes with binders such as bentonite, while Sludges are agglomerated into mixes. Those materials contaminate the products, are expensive and generate significant amounts of slag. Achieving cold dust agglomeration with little or no binder requires understanding the controlling parameters in order to predict and optimize the process behavior. This in turn necessitates comprehensive dust characterization to base predictions on measurable characteristics. In the present work we collected 13 Mn-rich dusts over all the process chain: feed fines, smelting, and refining dusts, as well as ferromanganese alloy dust. This is the first systematic study of dusts along the Mn-processing chain for chemical, physical and mineralogical parameters. X-Ray Diffraction, Qemscan® and laser granulometry analyses show that the mineralogy, Particle Size Distribution (PSD) and morphology evolves from complex heterogenous to simple, homogenous along the process chain: The mineralogy of the primary ore dusts shows various hydroxides, oxides, carbonates, tectosilicates and phyllosilicates. The furnace dusts contain oxides, tectosilicates and carbonates with volatiles, while the refining dusts present only oxides and metal. Finally, the Metal dust is nearly pure metal. Particle morphologies and PSD vary from jagged, irregular particles with broad PSD (>1 mm to <0.3 µm) to smooth and spherical with narrow PSD (100 µm to <0.3 µm). The Metal dust is an exception with angular particles. The Cationic Exchange Capacities (CECs) of all samples range from 0 to 22 meq/100g. Specific surface areas (SSAs) range from <0.1 to 27m²/g, and seem to decrease with from ore feed to Metal dust. This may be due to the disappearance of hydrated and/or sheet-bearing minerals, since a smaller grain size usually implies an increased SSA. The zeta potential varies strongly over the range of the samples tested. However, beside the Gabon ore which has an Isoelectric Point (IEP) ~ 4.1, the other IEP ranges between 2 and 2.5."

Jan 1, 2018

By Kenneth M. Smith

The kaolin deposits of Aiken County, South Carolina (fig. 1), have been worked continuously since 1852. It has been long known that the deposits contain a high percentage of&apos; alumina and that large reserves exist. These facts led to the early selection of&apos; this area for examination in the search by the Bureau of Mines for high-alumina clays. An exploratory program was started by the Bureau, and drilling was carried on from. June 27, 1942, to October 31, 1942. The kaolin-bearing district occupies an area approximately 25 miles long and 11 miles wide, running northeast, with its center at Aiken, S. C. (fig. 1). Within this district, a triangular area bounded by Horse Creek, Silver Bluff Road, and Hamburg Road was roughly prospected by boring 43 machine and 12 hand holes. As a result of the data obtained, a final smaller area was chosen for more systematic prospecting. This area is bounded by Pine Log, Tewkesbury, Silver Bluff, and Craig Sawmill Roads and covers about 4,000 acres. Mineral exploration agreements covering 11,272 acres were signed by 35 property owners, and 11,956 feet was drilled in 200 holes on these properties.

Jan 1, 1949

By J deVries, A Broodryk, P Kyselica, K McLean

Gwalia gold mine is located close to the town of Leonora in the Goldfields region of Western Australia, some 830 km from Perth. The mine is wholly owned by St Barbara Ltd, an Australian Securities Exchange listed gold producer and explorer. The Gwalia mine is a deep, all-trucking mine and is serviced by a 10 km long decline from surface. Mining is currently 1540 m below surface. Reserves extend to 1820 m below surface and are open at depth. The mine produces in excess of 200 koz of gold per annum.The Leonora area is hot and arid, with a mean annual rainfall of 230 mm and a 95 per cent mean daily maximum temperature of 20.3&#176;C wet-bulb and 35.6&#176;C dry-bulb.The primary ventilation system comprises a 490 kg/s surface exhaust fan, which draws air down a parallel intake system consisting of the decline, and a dedicated fresh-air intake system extending to the bottom of the production zone. Cooling is provided via a surface bulk air cooler serviced by an ammonia refrigeration plant that delivers 150 kg/s of chilled water to the bulk air cooler.The refrigeration plant and bulk air cooler were installed in 2008. The design allows cooling to be increased over time by installing an additional refrigeration plant in a lead lag configuration with the current refrigeration plant when required.Since 2008, expansion and increased depth reached a point where the underground heat load exceeded the installed capacity of the existing 4.0 megawatt rated (MWR) refrigeration plant. In order to maintain the mine design conditions underground, upgrading of the existing refrigeration system was required to bring the cooling capacity to 7.5 MWR.Two expansion options were considered: the addition of an ammonia vapour-compression refrigeration machine similar to the existing one or a lithium bromide absorption refrigeration system using waste heat from the site power station. The original design and layout of equipment on-site meant that either could be used, allowing an assessment of the merits of both when additional cooling was required.The decision was made to install a lithium bromide absorption system on the lead side of the existing ammonia refrigeration plant. This arrangement provides considerable economic benefits, including reductions in site-generated electrical power, generator fuel consumption and carbon dioxide emissions.This paper details the planning, design, installation and commissioning of the absorption chillers.CITATION:Broodryk, A, deVries, J, Kyselica, P and McLean, K, 2015. The design, installation and commissioning of an absorption refrigeration system at Gwalia gold mine in Western Australia, in Proceedings The Australian Mine Ventilation Conference, pp 193–200 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 31, 2015

By C. W. Sherman, John Chipman

IN the mathematical statement of the law of mass action, the activity of each substance consumed or produced in a reaction is used to obtain a numerical constant which is characteristic of the equilibrium of that reaction at the given temperature. The activity in many cases is equal to, or proportional to, the concentration. If this is the case, any units denoting concentrations, such as weight percentage or mol fraction, may be substituted in the equilibrium expression. It is even permissible to use mixed units provided only that each unit is that used in defining the activity of the corresponding substance. The indiscriminate use of concentration units in writing equilibrium constants sometimes leads to very erroneous results. On the other hand, what appear to be contradictory results from similar investigations often might be resolved if the activity is the quantity that is used to determine the equilibrium constant for the reaction in question. De-sulphurization serves as an important illustration. Sulphur in lron and Steel The control of the sulphur content of the finished product is one of the oldest and broadest problems in the metallurgy of iron and steel. In general, this control must be exercised through the action of a slag, and for this reason experimental studies of the distribution of sulphur between slag and metal have assumed considerable importance. Under slags of the type found in the basic open-hearth furnace the ratio (% S)/[% S] does not in general exceed about 8. Under blast-furnace conditions, however, de-sulphurization ratios of 50 to 100 are readily obtained, and under laboratory conditions using similar slags, the ratio may be as high as 400. The fact that blast-furnace metal is much more readily desulphurized than is molten steel rests upon two distinct phenomena. The first of these is connected with the slag reaction. In the open-hearth furnace, the equilibrium in the following reaction places a definite limit on the sulphur content of the slag: S + 0-- = 0 + S- - where the symbols underlined indicate elements dissolved in the liquid metal, while slag constituents are represented as ions. The high oxygen content of the open-hearth bath prevents this reaction from proceeding far enough to effect good desul-phurization. In the blast-furnace process, however, the oxygen activity is limited by the presence of carbon in the metal and in the coke bed and the controlling reaction becomes: S_+O- + C_=S- + CO(gas) [2] The second phenomenon which contributes to the easier desulphurization of blast-furnace metal is the effect of high concentrations of other elements. These elements, notably carbon, silicon, and phosphorus, increase the thermodynamic activity of sulphur several fold. As a result of this increase, the escaping tendency of sulphur from the metal phase is increased and its removal facilitated. Previous Studies It becomes, therefore, a matter of importance to determine the effects of various elements upon the activity of sulphur in the melt. The experimental approach to this problem has been made through the equilibrium between sulphur in the melt and gaseous atmospheres containing hydrogen and hydrogen sulphide. The reaction and its equilibrium constant are written as follows: S + H2(gas) =H2S(gas);K1 = -- [3] Pn2 • as where the symbol a, represents the activity of the sulphur in solution in the metal. The understanding of the chemical behavior of sulphur in steel and pig iron rests on a firm knowledge of the binary system Fe-S. A number of studies of the equilibrium have given closely agreeing results from which the activity and free energy of the dissolved sulphur may be found. The recent paper by Sherman, Elvander, and Chipman1 includes a discussion and quantitative comparison of the results of several of the significant researchesr-" on dilute solutions of sulphur in iron. The results of this study indicate that the relationship between the ratio Ph,s/p+ and the percentage of sulphur in solution is not a linear one. This

Jan 1, 1953

By Paloma Lazaro, Moe Momayez

Interactions between human beings and their work environment require the body to regulate its temperature by balancing heat production and loss. Comfortable environmental conditions are crucial for keeping workers safe and healthy and to maintain a suitable level of productivity. However, achieving a proper core body temperature may become challenging under different conditions, especially in high heat-generating workplaces such as hot underground mines. Because hot underground mines have the potential to expose workers to heat stress, compliance with standardized and regulated indices and criteria is distinctly required. The objective of this paper is to provide an overview of the current research on heat stress and strain in hot underground mines, collected from published and specialized literature. General definitions, statistics of heat-related accidents in mining, overview of the indices, standards, and recommendations for heat stress are provided herein.

Oct 6, 2020

By E. D. Gardner

FIFTY years ago the Utah Copper enterprise at Bingham was just getting under way. An epic in metal mining was in the making. Throughout the West the bonanza deposits were approaching exhaustion and most ore still went direct to smelter. But concentrators were shortly to be constructed for treating milling grade ores, and most important, the pilot plant for the Utah Copper enterprise was being built that year at the mouth of Bingham Canyon. The same decade was to see development at the Inspiration mine in Arizona of the most important new mining method for 50 years to come—undercut block caving. Although the average grade of ore being mined has been declining for 50 years, and wage rates and prices of supplies have been greatly increased, the mining industry has continued to grow and to prosper. Technology has advanced to a point where minerals can be mined under the most severe conditions. The hardest rock can be drilled and broken and the heaviest ground handled. Large or small, an ore-body can be mined successfully, at any desired rate of production. It is possible to pump great volumes of water and to ventilate the most extensive workings; air conditioning has made practicable the mining of ore where the rock temperatures are beyond those that can be borne by man. Mining today in the U. S. is a huge earth-moving operation. In 1953 nearly 2 billion tons of material were mined for direct use or for processing. Stripping at open-cut mines totaled about 1.7 billion cu yd. Table I gives relative tonnages of the branches of the mineral industry in 1953. In addition to the solid materials moved, 350 million tons of petroleum and 340,000 tons of gas were produced in 1953. Fifteen million tons of salt and 5 million tons of liquid sulphur were pumped from wells. An appreciable amount of copper was recovered by leaching old mine workings and dumps. Total value of minerals produced in 1953 was $14.4 billion: oil and gas $7.6 billion, coal $2.6 billion, nonmetallics $2.4 billion, and metals $1.8 billion. Formidable problems face some branches of the industry today, especially pertaining to increased output tons per manshift. The time has come when these problems should be reappraised. With current wage rates, the industry no longer can afford to pay men to do manual labor. It should be recognized, also, that the miner is a skilled worker; efforts to promote the dignity of his calling and his pride of accomplishment have yielded beneficial results. Mineral deposits occur under widely varying conditions. In most instances the degree of support needed by the roof and walls is the deciding factor in selection of an underground mining method. The degree of support needed, in turn, is governed by natural physical characteristics of the ore and its enclosing rocks. Mine operators, of course, learn to judge ground by observation, and an experienced mine boss generally can anticipate what will happen following any given mining operation. A more scientific approach, however, would be advantageous in most cases. More should be known about the strength of the ore and the enclosing rocks, more about stresses set up in the ground when ore is removed, and more about the mechanics of moving ground. Better understanding would work both ways: first, favorable rock conditions could be fully utilized for more efficient mining, and second, hazards of rock falls could be minimized, loss of ore in mining reduced, and expenses caused by ground pressures and subsidence lowered. In general, mining comprises breaking the ore or rock in place, supporting the mine workings, and transporting the ore to the surface. This latter is a materials handling problem. Ventilation, of course, must be provided, and the mines must be kept free of water. Development work at some places is the most important cost item in extracting ore. Drilling and blasting remain the usual practice for breaking ore or rock from place except, of course, at block-caving mines. The drilling and blasting practice at a mine must fit the mining method. In laying out mining procedures attention is being given to

Jan 1, 1956

By Frederick K. Allgaier

As part of an ongoing research program, the Bureau of Mines, U.S. Department of the Interior, is monitoring surface subsidence over six longwall panels at three Utah coal mines. These sites are characterized by mining depths of 244 to 610 meters (800 to 2,000 ft .), thick and massive sandstone layers in the overburden strata, multiple seam mining, extraction heights of approximately 1.8 to 3 meters (6 to 10 ft.), and rugged surface topography. Maximum subsidence, angle of draw, timing relative to mining, horizontal strain, and rate of subsidence development are being determined and correlated with mining and geologic variables. Results to date indicate the maximum vertical subsidence over the monitored panels ranges from 0.6 to 0.9 meters (2 to 3 ft.). The major factors controlling the amount of subsidence are the panel dimensions and the status of adjacent panel mining. Only minor surface cracking at the panel ends was evident at any of the sites. In all cases, there was a time lag of several months between the start of mining and subsidence at the surface. As subsequent panels are monitored and additional data become available, better estimates of the effect of various site characteristics on the subsidence will be possible. Ultimately, improved prediction capabilities will allow mine operators in the West to evaluate the potential for subsidence-related damage which could affect the current or future land use above the mine.

Jan 1, 1982

By David R. Kalasky

One of the primary areas of computer application in mine ventilation has been the analysis of mine ventilation networks. A variety of network analyzers that are aimed toward the solution of flows in a given network has been developed. These programs typically use the Hardy Cross or other iterative algorithm for solution of any network, even those that are complex. The computer software has become more efficient and sophisticated over time and has thus been widely adopted for ventilation network analysis. One of the shortcomings of these programs is that they have been developed with the idea of analyzing the ventilation requirements at a single point in the lifetime of a mine. This provides the ventilation engineer with an important tool for analysis but falls somewhat short of providing a convenient means to economically com- pare alternative methods for ventilating a mine. The program described in this paper provides such a tool. The principal concepts that were necessary for the completion of a computer program that analyzes ventilation alternatives is the capability of efficiently determining the ventilation requirements at a variety of times during the lifetime of the ventilation system and some method of analyzing the costs of the alternatives on a realistic comparative basis. To accomplish this task, the authors began using Penn State&apos;s mine ventilation simulator designated as PSU/MVS, Didyk et al. (1977a, 1977b) and added the necessary additional software to accommodate the need for multi-period, time-dependent analysis of the net- work. The software was then tested on a coal mine ventilation problem with a 23-year life. The problem points out the value of such a program for ventilation analysis.

Jan 1, 1986

By Thomas J. Fisher, John J. Sammarco, Gary L. Mowrey, Edward F. Fries

This report (Safety File 3.0) is the fourth in a nine-part series of recommendations addressing the functional safety of processor-controlled mining equipment. It is part of a risk-based system safety process encompassing hardware, software, humans, and the operating environment for the equipment’s life cycle. Figure 1 shows a safety framework containing these recommendations. The reports in this series address the various life cycle stages of inception, design, approval and certification, commissioning, operation, maintenance, and decommissioning. These recommendations were developed as a joint project between the National Institute for Occupational Safety and Health and the Mine Safety and Health Administration. They are intended for use by mining companies, original equipment manufacturers, and aftermarket suppliers to these mining companies. Users of these reports are expected to consider the set in total during the design cycle. • 1.0 Safety Introduction.—This is an introductory report for the general mining industry. It provides basic system/software safety concepts, discusses the need for mining to address the functional safety of programmable electronics (PE), and includes the benefits of implementing a system/software safety program. • 2.1 System Safety and 2.2 Software Safety.—These reports draw heavily from International Electrotechnical Commission (IEC) standard IEC 61508 [IEC 1998a,b,c,d,e,f,g] and other standards. The scope is “surface and underground safety-related mining systems employing embedded, networked, and nonnetworked programmable electronics.” System safety seeks to design safety into all phases of the entire system. Software is a subsystem; thus, software safety is a part of the system’s safety. • 3.0 Safety File.—This report contains the documentation that demonstrates the level of safety built into the system and identifies limitations for the system’s use and operation. In essence, it is a “proof of safety” that the system and its operation meet the appropriate level of safety for the intended application. It starts from the beginning of the design, is maintained during the full lifecycle of the system, and provides administrative support for the safety program of the full system. • 4.0 Safety Assessment.—The independent assessment of the safety file is addressed. It establishes consistent methods to determine the completeness and suitability of safety evidence and justifications. This assessment could be conducted by an independent third party. • 5.0 Safety Framework Guidance.—It is intended to supplement the safety framework reports with guidance providing users with additional information. The purpose is to assist users in applying the concepts presented. In other words, the safety framework is what needs to be done and the guidance is the how it can be done. The guidance information reinforces the concepts, describes various methodologies that can be used, and gives examples and references. It also gives information on the benefits and drawbacks of various methodologies. The guidance reports are not intended to promote a single methodology or to be an exhaustive treatment of the subject material. They provide information and references so that the user can more intelligently choose and implement the appropriate methodologies given the user’s application and capabilities. [ ]

Jan 5, 2002

By N. P. Peterson

The Castle Dome copper deposit is of the porphyry type and occurs in a body of quartz monzonite intruded into the pre-Cambrian formations and possibly into the lower part of Paleozoic limestones. The hypogene sulphide minerals include pyrite, chalcopyrite, and a little molybdenite, sphalerite, and galena. Pyrite and chalcopyrite are disseminated or occur in small closely spaced veins. The hypogene mineralization shows distinct zoning in the distribution of pyrite and chalcopyrite, so that the mineralized area can be subdivided into pyritc and chalcopyrite zones. Supergene enrichment, though not extensive, played an important part in the formation of the ore body, the chalcopyritc zone being the pro-tore. Several thin, gently dipping, diabase sills occur in the mineralized area, and the copper content of the rock in and near these sills is substantially higher than average. The formation of the ore body was caused by the combined effects of three forces: I. Zoning of the hypogene mineralization, 2. Richer copper metallization associated with the diabase sills, and 3. Supergene enrichment. Introduction The Castle Dome copper deposit is about 5 miles west of Miami, Arizona, and about 3 miles west of the Miami-Inspiration ore body. The Castle Dome open-pit mine is on the south flank of Porphyry Mountain, a prominent landmark of the region. In 1941 the deposit was being explored by the Miami Copper Co., but the exploration program was cut short when the decision was made to develop the property as a war project. The Castle Dome Copper Co., Inc., was organized as a wholly owned subsidiary of the Miami Copper Co., and arrangements were made with Defense Plant Corp. for developing and equipping the mine and for construction of a concentrating plant. Construction and development work was begun in January 1942, and milling operations started June so, 1943. Since that date, the mine has been in continuous operation by the Castle Dome Copper Co, under a lease agreement with Defense Plant Corp. producing copper at an average rate of 4,000,000 lb per month. To January I, 1946 the total production amounted to 121,088,000 lb of copper,4,271 oz of gold, and 211,755 oz of silver. A detailed study of the copper deposit and the surrounding area was undertaken by the U. S. Geological Survey in June 1943. Other members of the Survey participating in the work were C. M. Gilbert, G. L. Quick, and J. P. Albers. The results of this work will be published as a comprehensive report by the Geological Survey. General Geology The Globe quadrangle, which includes the Castle Dome area, was studied by F. L. Ransome1 of the Geological Survey in 1901 and 1902. His maps and descriptions were

Jan 1, 1949

By N. P. Peterson

The Castle Dome copper deposit is of the porphyry type and occurs in a body of quartz monzonite intruded into the pre-Cambrian formations and possibly into the lower part of Paleozoic limestones. The hypogene sulphide minerals include pyrite, chalcopyrite, and a little molybdenite, sphalerite, and galena. Pyrite and chalcopyrite are disseminated or occur in small closely spaced veins. The hypogene mineralization shows distinct zoning in the distribution of pyrite and chalcopyrite, so that the mineralized area can be subdivided into pyritc and chalcopyrite zones. Supergene enrichment, though not extensive, played an important part in the formation of the ore body, the chalcopyritc zone being the pro-tore. Several thin, gently dipping, diabase sills occur in the mineralized area, and the copper content of the rock in and near these sills is substantially higher than average. The formation of the ore body was caused by the combined effects of three forces: I. Zoning of the hypogene mineralization, 2. Richer copper metallization associated with the diabase sills, and 3. Supergene enrichment. Introduction The Castle Dome copper deposit is about 5 miles west of Miami, Arizona, and about 3 miles west of the Miami-Inspiration ore body. The Castle Dome open-pit mine is on the south flank of Porphyry Mountain, a prominent landmark of the region. In 1941 the deposit was being explored by the Miami Copper Co., but the exploration program was cut short when the decision was made to develop the property as a war project. The Castle Dome Copper Co., Inc., was organized as a wholly owned subsidiary of the Miami Copper Co., and arrangements were made with Defense Plant Corp. for developing and equipping the mine and for construction of a concentrating plant. Construction and development work was begun in January 1942, and milling operations started June so, 1943. Since that date, the mine has been in continuous operation by the Castle Dome Copper Co, under a lease agreement with Defense Plant Corp. producing copper at an average rate of 4,000,000 lb per month. To January I, 1946 the total production amounted to 121,088,000 lb of copper,4,271 oz of gold, and 211,755 oz of silver. A detailed study of the copper deposit and the surrounding area was undertaken by the U. S. Geological Survey in June 1943. Other members of the Survey participating in the work were C. M. Gilbert, G. L. Quick, and J. P. Albers. The results of this work will be published as a comprehensive report by the Geological Survey. General Geology The Globe quadrangle, which includes the Castle Dome area, was studied by F. L. Ransome1 of the Geological Survey in 1901 and 1902. His maps and descriptions were

Jan 1, 1949

By H. W. Imholz

This report covers most of the counties in North Central Texas in which oil is being produced in commercial quantities. The exceptions are McCulloch and Runnels, which were omitted because the production is small and the detailed information necessary for a report is not available to the writer. Complete information was not available regarding the number of wells drilled in this district during the year. The writer wishes to thank Mr. T. F. Petty, of the Humble Oil and Refining Co., Wichita Falls, Texas, for information used in compiling Table 2.

Jan 1, 1937