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By Madan M. Singh

With the current emphasis on ecology, safety, and conservation of natural resources, some consideration should be given to methods of mining that reduce surface damage, make the mining of coal less hazardous, and allow exploitation of a larger percentage of the coal in situ. One method that can be helpful toward attaining these objectives is the use of packing or stowing. Stowing involves the replacement of the ore that has been removed by waste material obtained from within the mine or elsewhere. Evidently, it would be most beneficial if this waste for packing were available from the mine being worked, but generally this is not possible. If complete stowing is adopted, it is necessary to bring in packing material from other sources. Stowing tends to reduce surface subsidence, but not eliminate it. It enhances roof support and curtails accumulation of methane in the gob, making the mining operations safer. It may also allow one to obtain a larger amount of coal. Currently, in the United States, a substantial amount of coal is left in place as pillars, in the hope that it will support the overburden indefinitely. Some of these could be extracted with stowing. The mining of thick, steep, or contiguous seams would also be facilitated.

Jan 1, 1974

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

"Underground refuge alternatives (RAs) require an air source to supply breathable air to the occupants. This requires pressure relief valves (PRVs) to prevent unsafe pressures from building up within the RA. The Mine Safety and Health Administration (MSHA) mandates that PRVs prevent pressure from exceeding 1.25 kPa (0.18 psi or 5 in H2O), or as specified by the manufacturer, above mine atmospheric pressure when a fan or compressor is used for the air supply. Research to date by the National Institute for Occupational Safety and Health (NIOSH) suggests that the design and implementation of PRVs used in RAs has not yet had sufficient performance analysis or technical development. In response to this need, NIOSH tested a variety of PRVs using an instrumented test fixture consisting of data acquisition equipment, a centrifugal blower, ductwork, and various sensors to determine if subject PRVs meet the MSHA requirement. Relief pressures and flow characteristics including opening pressure and flow rate were measured for five different PRVs under a variety of conditions. Subject PRVs included two off-the-shelf modified check valves, two check valves used in MSHA-approved built-in-place (BIP) RAs, and a commercially available valve that was designed for a steel RA and is currently being used in some BIP RAs. The test results showed relief pressures ranging from 0.20 to 1.53 kPa (0.8 to 6.1 in H2O) and flow rates up to 19.3 m3/min (683 SCFM). INTRODUCTION The Mine Improvement and New Emergency Response Act of 2006 (MINER Act) was enacted 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 [1]. Section 13 of the MINER Act specifically directed NIOSH to provide for research into the effectiveness and viability of refuge alternatives (RAs) for underground coal mines. This mandate culminated in the 2009 adoption of changes to 30 CFR mining health and safety regulations, requiring underground coal mines to provide mine emergency RAs and associated components, to provide a life-sustaining environment for persons trapped underground. Such RAs can be either self-contained mobile units or built-in-place (BIP) facilities. The regulatory changes also include provisions establishing requirements for Mine Safety and Health Administration (MSHA) approval of RAs and their components, and among these provisions are numerous criteria for providing a safe breathable atmosphere under positive pressure within the RAs. One specific criterion for maintaining a safe RA atmosphere requires the inclusion of an air pressure relief valve that will activate at a maximum of 1.25 kPa (0.18 psi or 5.0 in. H2O), or at a pressure above mine atmospheric pressure in the RA as specified by the manufacturer when the breathable air is supplied by a fan or compressor [2]."

Jan 1, 2017

By M. A. Anderson

After mope than fifty years Of Continuous operation, the Nevada Mines Division found itself faced with declining ore headings and metallurgical efficiencies, and rising operating costs. Although intensive process improvement programs had been initiated at various times in its history, insufficient return on investment served to defeat plans for any major changes in circuitry. The advent of practical large-volume flotation machinery, along with the success of long-range exploration drilling programs, complimented the results of an intensive program of research and development. The result of laboratory testing, pilot plant testing, and full scale plant tests was the proven indication of metallurgical improvement given increased flotation residence time and the effect of mechanical agitation in flotation.

Jan 1, 1975

By John Turner, John A. Gibson, David Cass

Contents A. Introduction B. Description of Robbins Company involvement in several of the English Channel Tunnel projects C. Identification of project requirements and comparison of TBM features D. Description of the manufacturers separate roles in the -235 (T-1) TBM E. T-1 performance to January 15, 1989 F. Summary

Jan 1, 1989

The United States is being invaded... by foreign capital: marks, francs, pounds, yen, lire, riyals, you name it: On almost any given day the Wall Street Journal will contain a new report on a department store, newspaper, bank, hotel, agricultural land, supermarket chain, et a1 being bought into by overseas investment. Xenophobic reaction to this revised national condition has begun to set in. Several farming states have and are introducing legislation to monitor foreign land buying. In Congress a mandatory reporting requirement for foreign purchasing is currently under study. In a relatively short period the country has reverted from world leader- ship in overseas acquisition and involvement to become itself one of the prime investment opportunities in existence. Interestingly enough, one of the industries where this changed status has become most apparent and pronounced is within that small/regional/segmented/low value-to-weight bulk commodity enterprise ranked as the backbone of the nation's construction activity. U.S. cement making is not only undergoing internationalization in its ownership, but also in its product, equipment, and engineering expertise.

Jan 1, 1979

By E. M. - Technical Papers Spokes

A.B. Rushton The work and resulting publications of the Committee on Underground Coal Mine Safety and The President's Commission on Coal were appropriate at the time of publication. The statistics cited in the publications appeared to be valid, and the actions proposed by the committee had merit. Although the actions proposed to achieve further improvements in safety may have been worthy of repeating, to associate statistics (Table 1) that are seven to eight years old with company names is bad. Statistics are readily available from the Mine Safety and Health Administration (MSHA) and are usually within three months of being up to date. A review of the most recent Coal Controlling Co. Report will reveal significant changes in the rankings. The more recent statistics should have been cited, or at least compared with those used in the article. Westmoreland Coal Co. is listed as having a 21.1 disabling injury rate for the period (1/78-12/80). The disabling injury rate for the Company for the year ending 1985 was 7.4 for the underground mines and 6.25 for the total facilities. Through March 1986, the dis- abling rate for the underground mines was 2.86. The management and employees of Westmoreland Coal Co. are committed to safety. The improvements achieved over the past five years attest to that commitment. We are justifiably proud of the achievement and dislike being misrepresented as in the April 1986 article Reply by E.M. Spokes I regret that the lapse of time since the original work gives a picture of safety practice that is no longer true today. That Westmoreland has reduced its disabling injury rate by 65% to 7.4 in 1985 is most commendable, and I congratulate you for this accomplishment. It is typical of the progress sought by the National Academy of Science and National Research Council when they funded the original study and published "Toward Safer Underground Coal Mines" in 1982. If a significant number of companies do as well as Westmoreland, the report will have served its purpose. Since the paper was primarily a condensation of the original report, it would not have been appropriate to seek access to later statistics. The purpose in seeking its publication at the time of presentation was to reach those who might have missed the original report. The fact that publication was delayed has not voided that purpose and may have reminded some who were aware of the original that its recommendations still apply to improvement of underground coal mine safety

Jan 1, 1987

By D. Malhotra, K. Purdy, G. Ramadorai

INTRODUCTION Today, flotation is clearly the dominant mineral concentration method. Since it is a relatively selective process, it is used for almost all sulfide minerals. A recent U. S. Bureau of Mines study indicated that approximately 225 million metric tons of sulfide ore was processed in 67 plants in the United States in 1980 (USBM Staff, 1981). There were over 100 plants processing copper ores in the world (Bassarear, 1985). Sulfide ores containing lead, zinc, copper, molybdenum, precious metals, cobalt, nickel, and iron are generally separated and concentrated by flotation. This chapter briefly reviews the plant practice related to the recovery of the above-mentioned sulfide minerals with emphasis on chemical reagents used for the production of marketable-grade products. PROCESSING OF SULFIDE ORES It is beyond the scope of this chapter to review the milling data for operating plants which have been extensively covered in literature (Tveter and McQuiston, 1962; Rausch and Mariacher, 1970; Weiss, 1985; Crozier, 1979). However, the following generalizations can be made regarding the processing of sulfide ores: 1. The primary grind is targeted to produce as coarse a product as necessary to provide an economic degree of liberation. 2. The concentration process, generally flotation, is designed to produce a concentrate consisting predominantly of valuable minerals and a middling fraction which may require further grinding to produce an optimum release of minerals. 3. The chemical reagent suite (collectors, frothers, etc.) for the separation of sulfide minerals from gangue minerals is generally complex. Total reagents used in the flotation process can be as high as six to eight, as is the case for lead and molybdenite separation from complex sulfide ores. 4. Chemical reagents are generally expensive because of the high selectivity for sulfide minerals. The process flowsheet is dependent on several factors, including ore grade and mineralogical composition. Generic process flowsheets for processing different sulfide ores have been developed for discussion in this chapter. However, there appears to be a consistent relationship between the reagent combination choice and the mineralogical composition of the ore (Tveter and McQuiston, 1962). These are covered in detail in the following sections. LEAD AND ZINC Galena (PbS) and sphalerite (ZnS), which commonly occur together, are the only minerals of lead and zinc, respectively, in most primary deposits. Feed ore grades are typically 1 to 5 percent Pb and 1 to 10 percent Zn. Very often economic quantities of copper and precious metals, silver and gold, also occur with these base metals.

Jan 1, 1986

By A. J. Benham, H. Kessler

In the UK a high proportion of land-won sand and gravel is extracted from sub-alluvial gravels and river terrace gravels located within river valleys. However, extraction from deposits of this type is increasingly proving to be less acceptable. Quarries within river valleys often have a high visual impact and the cumulative impact of these operations in close proximity to each other can be high (figure 1). River valley operations are almost always restored to open water. Whilst these may be attractive for nature conservation and provide leisure activities, they can also have a serious impact on the natural hydrology of the flood plain. Wet restoration of quarries located close to airports or airfields can also lead to concerns regarding increased risk of bird strike on aircraft. The Civil Aviation Authority recently suggested that an exclusion zone of 13 km be applied around all airports and airfields to reduce the potential for bird strikes on aircraft. This action is likely to have a serious impact on the availability of potential sand and gravel extraction sites in river valleys.

Jan 1, 2005

By Louis W. Cope

INTRODUCTION Various methods for sampling placer deposits have been mentioned by Wojcik in Chapter 2. Here, a brief description and comparison of these methods will be given as related to bulk versus drill sampling. The advantages and disadvantages inherent to each will also be given. The means of sampling can be grouped into bulk or drill samples. As a generality, drill samples over a five-foot length will run from I/,, to 1 /, cubic yard, depending on the diameter of the drive shoe on the drill casing. Bulk samples can be expected to be in the range from one to many cubic yards each. SAMPLE METHODS Bulk Bulldozer-dug trenches and dozer or front-end loader pits are two methods to excavate large bulk samples. Smaller bulk samples can be obtained by digging small dimension shafts through the deposit, by hand or caisson-like drills. It is not recommended to split samples by alternate bucket-loads or by other means, as the effect of a nugget falling into one or the other fraction will seriously alter the sample.

Jan 1, 1992

By [Miroslaw] M. Hajdasinski

The Payback Period (PBP) is one of the most widely applied project evaluation criteria 'in the mineral industry. As much as this criterion is liked by practitioners, it is also criticized by academicians who seriously question its validity. Inspired by this criticism, the author of this paper examines the classical definition of the PBP criterion, and formulates a new one designed for both investment and borrowing projects. It is demonstrated that the PBP, in its revised version, is free from all the shortcommings of the old one, and is fully compatible with the Net Present Value (NPV) criterion, as well as with all the other NPV compatible project evaluation criteria. Since the traditional approach to the comparison of mutually exclusive projects (MEP1s) by means of the PBP criterion has been incorrect, the paper also outlines the correct approach to this problem. Finally, owing to the fact that the PBP depends on the assumed interest rate, and using a specially developed computer program, the PBP is graphically presented, in six numerical examples, as a function of the interest rate.

Jan 1, 1989

By Kim A. Lapakko, David A. Antonson, Jon R. Wagner

This study examined the potential of mixing rotary kiln fines (RK fines), a waste product generated when limestone is converted to lime (CaO), with acid producing mine waste to provide 1) short-term mitigation by neutralizing acid produced by iron sulfide oxidation and 2) long-term mitigation by facilitating the passivation of iron sulfide mineral surfaces. Five loadings of RK fines were mixed with 75 g of Duluth Complex rock to produce neutralization potential (NP) added: acid potential (AP) quotients of 0,0.11,0.22,0.66 and 1.1. The mixtures were subjected to wet-dry cycling laboratory testing for 585 weeks. The rock alone produced acidic drainage, with pH decreasing below 4 after about 20 weeks of reaction and ultimately reaching the approximate range of 3.3 to 3.5; sulfate release did not decrease over time. The RK fines elevated pH and inhibited sulfate release for time periods which increased with the mass of RK fines present. The pH of drainage from the 0.11 quotient (NP added1AP) loading decreased below 6.0 after 75 weeks of dissolution and decreased to 3.6 after 117 weeks, at which time the reactor was terminated. The duplicate 0.22 quotient loadings of RK fines decreased below 6.0 after 95 and 170 weeks. The former was terminated after 117 weeks, and the latter reached a minimum pH of 2.8 after 512 weeks. Drainage pH from the 0.66 quotient decreased below 6 after 581 weeks. For the drainages which acidified, the cumulative calcium release prior to pH decreasing below 6.0 was 92 to 103 percent of that initially present as calcium carbonateloxide in the RK fines. The pH of drainage from the 1.1 quotient typically remained above 8.0 throughout the period of record. Sulfate concentrations in the drainages decreased while the neutralizing fraction of the RK fines was present, but increased after this fraction was depleted.

Jan 1, 2000

By A. W. Khair, S. S. Peng

Background The mine is located in Ebensburg, Cambria County, PA. Cambria County is located within the Allegheny Mountain section of the Appalachian Plateau Province. The geologic structures in the mine area included folds, faults, joints and cleats in the coal. These geologic structures seem to have genetic relationships with the Allegheny Front (lannacchione and Puglio, 1979). The mine's geologic structures have been influenced mainly by Ebensburg Anticline and Wilmore Syncline (Watson, 1979). Current longwall panels lay along the western limb of the Wilmore Syncline oriented south 30° west. Surface topographic relief in the area is less than 152 m (500 ft) and overburden thickness varies from 183 to 244 m (600 ft to 800 ft). The coal seam in the mine is part of the Lower Kittanning Formation. The thickness of the B-seam is the subject of discussion. It varies from 1.2 to 1.5 m (48 to 60 in.) and consists of two parts. The lower part is 1.1 to 1.2 m (44 to 48 in.) of soft coal and the remainder is bony coal. The coal seam dips slightly from 0° to 6° to the east-south direction. Panel entries are driven 61 m (200 ft) wide, consisting of three 5.5 m (18 ft) wide entries and two pillars 24.4 x 30.5 m (80 x 100 ft) centers. The mine has a history of cutter roof problems in the entries (Fig. 1) that has delayed the advance rate of entry development considerably and the cost of maintaining those entries is very high due to the requirement of very heavy artificial supports (Fig. 2). Due to excessive convergence in the headgate, installation of additional support and often resetting of steel beam, to provide enough clearance, in these entries imposed additional delay during the retreat of the longwall face, as the shearer cuts coal adjacent to the headgate and shield support at the headgate needs to advance under the steel beams (Fig. 3). Methodology Extensive underground observation indicated that roof cutter [ ]

Jan 1, 1992

By Frederik P. Agterberg

Mineral exploration consists of the acquisition of data and their interpretation to ascertain the presence or absence of uncertain events. In general, the amount of uncertainty decreases during exploration when targets become increasingly well defined. Decisions on further exploration and whether or not mining should be commenced are based on new information using both objective facts and subjective reasoning. Computers play an important role during all stages of exploration. Three recent books (References 1-3) on decision-making in exploration and mineral-resources evaluation provide additional background supplementing the papers in this volume and articles in scientific journals or other conference proceedings (e.g., References 4 and 5) dealing with the subject. Other new developments to be considered are that microcomputers are becoming increasingly important in the earth sciences (Reference 6), and so are applications of artificial intelligence (Reference 7). Microcomputer- based technology allows rapid communication between field parties and main office. Expert systems are becoming increasingly sophisticated aids for mineral exploration an can now contain objective rules based on statistical theory (Reference 8).

Jan 1, 1989

By Scott A. Norby

Kennecott's Ray Mines Division (RMD) is located in central Arizona. The pit and silicate plants are located at Ray while the 27,500 tpd concentrating facilities are located at Hayden, approximately 20 miles away from the pit. Ore is delivered in 100 ton railroad cars, 60 cars per train via the Copper Basin Railroad. The side discharge cars are dumped into a 6,000 ton trackhopper. There are 6 Nico feeders, two per belt. Three 48" belts transport the -8" ore to three 7' standard (secondary) Nordberg crushers. Each crusher is preceded by a 6' x 12' Tyler double deck screen. The secondary crushers are open circuit. The -1/21t material from the screens is transported directly to the fine ore bin at the mill. The discharge from the crushers and the intermediate product -2 1/411 x +1/2" is conveyed to a surge bin. The tertiary crusher feeder belts discharge onto five 61 x 161 Tyler double deck screens, each followed by a 7' shorthead (tertiary) Nordberg crusher in closed circuit. The -1/2" product joins the -1/2" ore from the secondary crushers and both are conveyed along a 42", 4,230 foot long belt to the fine ore bin that feeds the grinding circuits. When hard ore is experienced at the concentrator the daily throughput decreases by up to 10%. This coupled with the fact that the ore is generally becoming harder with mine depth provided the incentive to improve the performance at the crushing plant.

Jan 1, 1986

By Madan M. Singh

Section 516 of Public Law (PL) 95-87, commonly known as the Surface Mining Control and Reclamation Act of 1977 (SMCRA), requires that underground coal mine operators must "adopt measures consistent with known technology in order to prevent subsidence causing material damage to the extent technologically and economically feasible, maximize mine stability, and maintain the value and reasonable foreseeable use of such surface lands, except in those instances where the mining technology used requires planned subsidence in a predictable and controlled manner..." Implied in the constraints of Section 516(b)(1) of SMCRA is a determination of what constitutes "material damage" to surface structures, renewable resource lands, and the capacity for sequential land use following underground mining. The authors have already briefly addressed the subject of damage criteria in relation to surface structures (Bhattacharya et al, 1984). In this paper, an attempt is made to provide quantitative guidelines to ascertain material damage to renewable resource lands, although the data available to establish these criteria are considerably more limited than is the case for surface structures. Besides, many more variables influence the severity of the impact at a given site, making the task of assigning stringent ground movement limits even more complex. This often necessitated use of professional judgement and occasional reverting to a qualitative approach. Hence the reader is cautioned at the very outset that the quantitative limits presented herein are approximate and are primarily intended to serve as rough guides for ascertaining the extent of damage to the renewable resource lands under consideration, and that depending on site-specific conditions the actual damage incurred may differ significantly from that predicted by the proposed criteria. The authors, however, feel that the criteria suggested are somewhat conservative, but realistic. The actual damage observed should generally be less severe than predicted by these criteria. It may also be clarified here that the term "subsidence", as used in this paper, implies the total phenomenon of surface and subsurface effects associated with the mining of minerals and not only the vertical displacement of the surface as is sometimes inferred in the literature.

Jan 1, 1984

By Ricardo D. Presnell, Erich U. Petersen, William J. Tafuri, Donald M. Hausen, Russell C. Babcock, William L. Gunter, Thomas R. Gibson, Douglas N. Halbe, Jay W. Hammitt

The Barneys Canyon orebody is a sediment-hosted disseminated gold deposit located 8 km (5 miles) north of the Bingham Canyon porphyry copper mine. Mineable oxide reserves are 8.6 million tonnes (9.5 million short tons) grading 1.64 g/tonne (0.048 oz/ton) gold. Host lithologies are carbonates and siliciclastics of the Park City and Kirkman-Diamond Creek Formations of Lower Permian age. Gold distributions are controlled by favorable stratigraphic units and associated alteration, and structures. Dominant alteration is decalcification and bleaching of carbonates, and argillization and bleaching of sandstones. Gold to silver ratios are greater than 1:1. Anomalous trace metals include arsenic, antimony, mercury and thallium. The Melco deposit lies 2.4 km (1.5 miles) southwest of the Barneys Canyon deposit. It is spatially associated with a steeply northwest-dipping siliciclastic breccia in the Kirkman-Diamond Creek Formation at the contact with underlying orthoquartzites of the Freeman Peak Formation. Both gold-bearing oxidized and sulfide lithologies are present, often in sharp contact. Mineable oxide reserves are 2.8 million tonnes (3.1 million short tons) grading 2.40 g/tonne (0.07 oz/ton) gold. Anomolous trace metals include arsenic, antimony, mercury and thallium, which are generally present in concentrations greater than those at Barneys Canyon, particularly in sulfide-bearing material. A spatial, temporal and genetic relationship is proposed between the Melco deposit and late-stage gold mineralization at Bingham Canyon, although both the Melco and Barneys Canyon deposits have many similarities to Carlin-type gold deposits.

Jan 1, 1990

By Deepak Malhotra

The first step in the development of a gold deposit is the detailed mineralogical examination of the ore types, major and minor mineral constituents and their association. It is extremely difficult and expensive mineralogically to characterize the occurrence and association of gold with other minerals for low-grade refractory ores (less than 0.1 troy oz/t Au). Even if a mineralogist can identify a gold-associated particle or two, one wonders how representative they are of the deposit. What alterative does a metallurgist have for characterization of gold-bearing ores? Diagnostic leach procedures, consisting of a series of sequential leaches, can provide the metallurgist/mineralogist with reasonably good Indication of the deportation of gold in the various mineral matrices. This information con help the metallurgist concentrate on the most promising process options In the Initial stages of the project development process.

Jan 1, 1993

By J de Cuyper, E. J. Parkins, H. L. Shergold, W. Freyberger

The paper is divided into two parts. The first part describes an investigation of the effect of temperature on the conditioning and flotation of an ilmenite ore with a 1:l oleic acid-fuel oil mixture as collector. During conditioning the pulp underwent characteristic changes of flocculation, deflocculation and dispersion which were monitored by measuring the power consumption of the conditioner motor. Optimum flotation results were obtained at the end of the deflocculation period. High ilmenite recoveries were obtained at high temperatures and high concentrate grades at low temperatures. The second part of the paper describes a study of the effect of temperature on the adsorption of oleic acid at the ilmenite/water interface. At high adsorption densities it was chemical and exothermic. Contact angles showed that the hydrophobicity of the collector-coated minerals increased with an increase in temperature.

Jan 1, 1976

By Wayne A. Sadik

The mine ventilation system of the Henderson Mine provides a case study of designing and maintaining high-quality respirable air in a large underground mine with over 200 units of diesel-powered equipment. This system was designed with concern for minimizing exposure to personnel of siliceous dust, diesel exhaust, smoke, and potential fire hazards while optimizing general environmental air quality throughout the multilevel/multiheading operation. As a relatively new operation, the Henderson Mine incorporates a modern ventilating plant and mine de- sign as well as state-of-the-art technology where appropriate. Two main ventilation shafts approximately 2,300 ft (701 m) in depth (one intake, one exhaust) will handle 1.8 million cfm (850 m3/s) of air circulated throughout 60 miles (97 km) of drifting, providing the ventilation requirements for more than 25,000 diesel horsepower (18,642 kw) underground.

Jan 1, 1980

By M. Schultz

The cost of remediation often drives the selection of the preferred remedial alternative at hazardous waste sites. The use of probabilistic cost/risk analyses can aid in predicting costs of remedial alternatives. Three types of information are discussed in performing cost/risk analyses: (1) data on the quantity of contamination versus the contaminant concentration; (2) unit costs of remediation for the remedial alternatives; and (3) risk to contaminant concentration relationships. Cost/risk evaluations are performed on a site-by-site basis, to account for combined risks of different contaminants and differences in quantity to be remediated versus contaminant concentration. The cost/risk analysis produces a graph of risk versus cost for each of the remedial alternatives. These costs are then compared to facilitate selection of a preferred alternative. An example of comparison of alternatives is illustrated for a historic coal gasification site.

Jan 1, 1992