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By R. H. King

Recently, the North American mining industry experienced a severe recession, forcing managers to take dramatic steps to cut costs and compete in the difficult international market. Some of these steps were closing mines, reducing work forces, renegotiating wage agreements, and purchasing the most productive equipment. Managers are now looking beyond these traditional avenues and are focusing more on advanced technology. In the present environment, it is essential that mine operators obtain the maximum use of capital expended on equipment. However, mine workers do not obtain maximum efficiency or productivity from equipment because the adverse and hazardous mine environment impedes human performance. Also, the efficient use of large, complex machines calls for levels of precision that many times are beyond the capability of even highly trained miners. A possible solution to this problem is a machine that mines without operators - autonomous mining machines. However, numerous problems confront researchers who are attempting to develop such equipment. Some mining tasks are not always composed of a series of cyclic motions readily performed by factory floor robots. In addition, mining takes place in the geological environment where conditions are highly variable and unpredictable. As a result, these machines must be able to sense and adapt to variations in operating tasks and environment. Considerable autonomous vehicle research has been completed, especially for defense. However, autonomous vehicle (AV) technology has not advanced to practical application yet (King, 1988) and mining research is necessary in areas: representing mining specific knowledge; •analyzing and reasoning about sensor data in the mining environment; and •discovering completely new mining methods or new approaches to existing methods that become apparent when we remove the constraints imposed by the necessity of human operators. A specific machine, the LHD, can be used to show the problems for researchers who are attempting to develop autonomous machines for mining. This author chose the LHD because it can borrow concepts developed for autonomous navigation by military programs. But considerable mining research is also required. Furthermore, studies done at the Henderson mine, in Colorado, show autonomous LHD 's promise cost and safety benefits (King, 1988). At Henderson, LHDs load ore from draw points, tram to an ore pass, dump and return, or switch to another draw point. An autonomous LHD must sense vehicle position along the route, relate sensor data to stored-map information, to determine location, follow drift center lines, plan paths between dump position and initiate appropriate control commands, sense vehicle operating status and vehicle health, key on features or targets for special tasks like high speed turns, perform end of travel tasks (loading and dumping), and detect and avoid obstacles. These goals are similar to those for shuttle cars, trucks, and front-end loaders. Therefore, much of the technology is transferable. Each Henderson LHD extracts six to 40 dippers from each of a series of draw points. The LHD transports the ore to an ore pass within 55 m (180 ft) of the draw point, making the longest round trip 110 m (360 ft). The LHDs have very fast hydraulic dumping and loading systems that reduce the round trip cycle to less than one minute. Even though the LHD is capable of 500 trips per shift, the average production is 300 dippers. Man trip and lunch reduce available operating time to 6.5 hours per shift. Mucking difficulties (setting large boulders aside), operator breaks for activities like talking shop, and cleaning and smoothing roads further reduce operating time. Supervised autonomy can reduce the number of operating units by increasing operating time per shift since computer controlled machines can operate during lunch and between shifts and reduce operator errors. If dippers per shift increase from 300 to 350 (long-range goals are 500 dippers per shift), constant production requires only 10 operating LHDs and two spares. Manpower requirements drop from 24 to four by controlling five machines from one workstation. A review of technology available from the Autonomous Land Vehicle, the Advanced Ground Vehicle Technology, the Ground Surveillance Robot, and other programs, show the following differences between others work and mining industry needs: •If we focus initially on mobile haulage vehicles, we can navigate from a map. We do not need to explore. •We can modify the environment to reduce the navigation requirements. •We have a harsher environment than any of the research programs have encountered. •Our equipment must operate faster and more precisely than present AVs. •Our equipment must operate reliably over long periods of time. •We must have better onboard machine health monitoring and diagnostics. •The AV programs do not address geosensing. •The major AV programs are not cost-effective. For example, we cannot afford the computer power for robust image processing, yet. To computer control an LHD, we must replace the guidance and monitoring skills of experienced operators with sensors, computing hardware, interfaces, and several software mod¬ules. Experienced operators avoid collisions and load efficiently in piles that may contain oversize muck. Collision avoidance without an operator requires sensing all obstacles in the draw

Jan 1, 1991

By M. Zaske, E. R. Mortensen

"As mine operators continuously look for ways to improve functionality and increase capacity of their facilities, existing process equipment is often pushed to and past its designed limitations. As a result, equipment manufacturers are asked for economical solutions to meet redefined processing needs. Newmont USA (Newmont) in Nevada was faced with such a need in their Carbon-in-Leach (CIL) train. In a collaborative effort between Newmont and Derrick Corporation, tests were conceived to assess the viability of drop in replacement screen using a three dimensional screen panel. The urethane pyramid screen panel was designed to increase flux while maintaining the existing equipment footprint. The test is the end result of an ongoing research and development project that Derrick undertook in an effort to refine and upgrade their urethane screening media. The application of this new screening media is presented for this CIL application along with future plans to implement the technology in trash, safety, carbon dewatering, carbon sizing, and other scalping applications.INTRODUCTION In Spring of 2014, Process Machinery Associates (PMA), Derrick Corporation’s (Derrick) local area representative, met with Newmont USA (Newmont) to discuss the operation of the in tank screens in the carbon in leach (CIL) train at Twin Creeks, near Golconda, Nevada. The process was originally designed for 13,000 +tons per day (tpd) but has been steadily rising since the plant was commissioned and now Newmont requires 16,000 tpd to be processed. The original screens were running at their maximum capacity and Newmont required additional screening capacity to continue the upward trend in production. Physical separation by use of screening is one of mineral processing oldest methods. Screening media can be prone to blinding and pegging and the necessary support structures required for finer apertures result in smaller open areas. Therefore, more screening area generally requires larger panels and therefore larger machines, screening media that doesn’t blind, or a combination of both. In an effort to avoid larger machines, Derrick Corporation had been working on a revolutionary three dimensional panel which would add non blinding screen area in the same footprint of an existing machine. Alpha testing of the new screen panels began at Newmont Mill 2/5 (Carlin, Nevada) in July of 2013 and was still occurring when PMA approached Twin Creeks about beta testing a refined machine and pyramid panel. After presenting the initial results, Twin Creeks determined that an upgraded machine and pyramid panels would be a possible solution to meet production needs. The testing of a machine eliminated the risk of capital costs to Newmont and provided Derrick the opportunity to obtain additional data."

Jan 1, 2015

By Shunji Sukizaki

Recently development of cheaper and cleaner energy sources has become a more important and urgent issue, because of the soaring crude oil price and environmental problems such as global warming. In this social context, to establish basic technologies and solve fundamental issues associated with environmental impact assessments, the Environmental Impact Assessment Group of the Research Consortium for Methane Hydrate Resources (MH21 research consortium) in Japan has conducted a number of research and development activities since FY2001. These activities include: 1) surveys of the marine environment that will form the basis of environmental impact assessments associated with methane hydrate development; 2) the development of core technologies to monitor deep-sea environments; 3) research of major issues related to environmental impact assessments; and 4) the development of numerical models to predict the deformation of seafloor sediments containing methane hydrates.

Jan 1, 2008

By R. A. Kirkland

The oxygen consumption method utilises an oxygen sensor to monitor the gas-phase oxygen concentrations in sealed chambers filled with reactive waste rock with subsequent confirmation of the oxygen content by gas chromatography. Results of experiments utilising low sulphide (<0.5w1%) waste rock yielded reproducible oxygen consumption rates on the order of 10-10 mol O2.kg-1. s-1. Petrographic analysis of the waste rock material indicated that the dominant sulphide mineral is pyrite (FeS2). Assuming a stoichiometry of 1.75 moles of O2 per mole of SO4 sulphate release rates were inferred to be between 6 and 18 mg SO4 kg-1 wk-1 Typically these results are 2 to 3 times greater than measured sulphate release rates from humidity cells. Our findings suggest that this techniques offers a rapid. cost efficient method for determining sulphate release rates in waste rock piles, even at low sulphide concentrations.

Jan 1, 2000

By CLAIRE STREBINGER, ADITYA JUGANDA, JURGEN F. BRUNE, GREGORY E. BOGIN JR.

Methane gas explosions are a major risk in underground coal mining operations. The severity of such explosions can range from local area damage to massive loss of miners’ lives along with significant damage to mine infrastructure and ventilation controls that may lead to mine closure and loss of the entire operation. This study demonstrates the viability of integrating a computational fluid dynamics (CFD) combustion model into a longwall mine ventilation model to simulate methane gas explosions in the longwall face area. The resulting 3D methane gas explosion simulation can provide better understanding of the fundamental physics and the potential impact of an explosion in an underground longwall coal mine.

Aug 1, 2022

By David E. Lindorff

The impact of surface coal mining on the geologic materials and ground water in different geologic settings is being investigated by the Illinois State. Geological Survey at three abandoned surface mines in Illinois. At a site near Ottawa in La Salle County, the overburden prior to mining consisted largely of pyritic shale; sand was the major component of the overburden and spoil at a mine near Coal City, Grundy County. Clays, lacustrine silts, and some shale and limestone were the dominant premining overburden materials at a mine near Harrisburg, in Saline County. Ground-water elevations at Coal City and Harrisburg are approximately at premining levels, whereas at Ottawa they are near the base of the spoil. Ground-water quality has deteriorated at all three sites, especially at Ottawa and Harrisburg.

Jan 1, 1981

By Dennis Dolinar

Improperly designed support systems have led to unplanned roof falls in longwall tailgates. To prevent such falls requires appropriate support and adequately designed support systems. The National Institute for Occupational Safety and Health (NIOSH) is currently developing a standing support design methodology based on the ground reaction curve. To design the support system with this methodology requires a quantitative assessment of the ground and support interaction, which is developed from instrumented test sites installed in longwall tailgates. To compliment and expand the field data results and to optimize the support design, numerical modeling is also conducted. The models are calibrated from the test site data. In the present study, data was obtained from a longwall tailgate in the Illinois basin. The test site was located in the Herrin No. 6 coal seam at a depth of 500 ft. At the site, the immediate roof was a weak shale overlain by a competent limestone. Above the limestone, sandstone formed the intermediate roof. The floor consisted of a weak underclay. In the study, the performance and interaction with the surrounding rock mass of two different standing supports, an engineered crib and a conventional 4-point wood crib were monitored. The two cribs have similar load-displacement characteristics at least through 6 in of displacement. Beyond 6 in, the engineered crib becomes unstable and begins to strain soften and shed load. The measurements and developed ground reaction curves indicated that the tailgate at the mine was a low convergence environment until well inby the face. Because the convergence was well below their capacity, both supports provide more than adequate levels of support to the tailgate. Significant convergence and support loading did not occur until 90 to 100 ft inby the face. The results of this study were also compared with results from similar studies conducted in the Pittsburgh seam to further advance the development of the ?ground reaction curve? based support system design.

Jan 1, 2009

By Floyd C. Bossard

A method of hand-calculating the air-cooling requirements of a conceptualized underground mining operation is presented for the reader&apos;s orientation. Separate air heat load calculations were conducted for adiabatic compression, electromechanical equipment, wallrock, broken rock, groundwater, and blasting operations. The air heat sources were calculated for four mining levels under conditions representative of a typical planned mining operation at depth. The total heat gain on each level will approximate the level air-cooling requirements. The results of these hand calculations can be further modified by the use of mine ventilation computer software that refine the heat-source calculations, predict underground ambient air temperatures, and establish the air-cooling requirements of a mine. INTRODUCTION The principal mining method employed is a modified vertical crater retreat (VCR) blasthole operation. Typical scopes range from 20 to 40 feet wide, 75 to 100 feet long, and 100 feet in height. Six and one-half inch blastholes are drilled with an "in the hole" hammer drill. Eight-foot high horizontal rounds are blasted down. Mucking of ore from the undercut to orepass is done with LHD equipment operated from a remote control station. Backfill includes hydraulically placed tailings with cement, and waste rock when available. Stope access is from ramp sub-levels on 50-foot vertical intervals. Crosscuts are ramped down to the first stope cut 25 feet below the sub-level elevation. Then the crosscuts are raised by taking down the back when each stope cut is completed, until an elevation 25 feet above the sub-level is reached. The crosscuts are filled with tailings, and/or waste rock. See Figure 1. Typically, a two-pronged approach to defining the air-cooling requirements is conducted. First, the principal sources that make up the air heat load are individually hand-calculated. Second, projections of mine heat load are calculated by utilizing computer modeling techniques. This paper discusses the first method (hand calculations) of determining the individual components of heat flow into the mine. CALCULATED AIR HEAT LOADS Adiabatic Compression The plans for the mine include delivery of 300,000 cfm of air to ventilate the lower level operation. This is equivalent to approximately 200 cfm/ton of rock produced (300,000 cfm/1500 tpd of ore and waste). As air flows down a shaft, with no heat interchange between the shaft and air and no evaporation of moisture taking place, the air is heated in the same way as if it were compressed in a compressor. Dry air increases in temperature about 5.4°F per 1000 feet. One BTU is added to each pound of air for every 778 feet of decrease in elevation, or is subtracted for the same elevation increase. For dry air, the dry-bulb temperature change is 1/(0.24 x 778) = 0.00535°F/ft., or 1°F/187 ft. elevation. Auto-compression may be masked by the presence of other heating or cooling sources, such as shaft wallrock, groundwater, air and water lines, electrical facilities, etc. The major factors influencing the temperature of the air delivered underground by a shaft are (1) the night time cool air temperature&apos;s effect on the rock or lining of the shaft, (2) temperature gradient of ground rock related to depth, and (3) evaporation of moisture within the shaft which increases the latent temperature and decreases the sensible temperature. [Calculation For Adiabatic Heat of Compression] a. Assumptions: 1. Three hundred thousand cfm of fresh air at 3000 level has increased in temperature during the summertime to the point where it has little available cooling power. Air-cooling will be required on 3000 Level and below. 2. Elevation of 3000 Level is approximately +1000 ft. above sea level.

Jan 1, 1993

By W. E. Horst

Historically, research organizations have made significant contributions to the development and commercialization of new chemical products for the mining industry and examples of commercial results are cited With the recent demise of research organizations within the mining industry and dynamic changes in the industry, the supporting role of R&D groups has become limited or non-existent. Cat- reduction requirements and new raw-material sources offer future opportunities for R&D organizations to continue their effective participation in the commerical development of new chemical products. Potential areas of interest for future process applications are reviewed.

Jan 1, 1986

By Mollie Hunt, William Riley, Russ Downer, Kate Willa Brown Requist, Víctor O. Tenorio, Geoffrey Gill

One of the first activities after returning to the Moon will be finding sources of water and materials for the construction of human settlements and other infrastructure. There is evidence of the existence of deposits in the bottom of craters at the South Polar region and in flat areas around. Based on geological data from the Moon, a block model representing a deposit or icy regolith has been defined in the vicinities of Shackleton Crater, with unique characteristics of shape, thickness, and water contents. A mine plan utilizing surface mine modeling software for ore extraction from an open pit- type of excavation to a delivery point near a water processing facility is proposed, with monthly, quarterly, and yearly schedules. Production results, dashboard charts and progress of the topographic changes are also presented.

Feb 1, 2024

By D. M. Ziebarth

"During the mid to late 1960’s, the underground potash industry in Saskatchewan was born. For the next four decades the industry operated at a steady pace supporting a large part of the world demand; until 2008 when the potash market heightened. Operators were motivated to capitalize on this increased demand, new mines were proposed while existing mines were planning upgrades. Potash Corporation of Saskatchewan (PCS) was one of the major producers planning upgrades to their production facilities quickly, including modifying the existing headframe and shaft at the PCS Cory Site to accommodate a new 1500 tonne per hour hoisting plant. Upon completion of the project, the innovative operations group continuously reviewed methods to increase production and reduce operating costs, allowing PCS to continue being competitive on the world market. This paper will describe the design challenges encountered during the development of the infrastructure around an operating facility whilst minimizing down time, schedule, capital and construction costs. This paper will also explore the strategy to increase capacity of the newly commissioned hoisting plant. BACKGROUND Potash production is a major Saskatchewan industry, and has played a significant role in the economy for over 40 years. Production commenced in 1962 when water problems were overcome at the International Minerals and Chemical (IMC) mine near Esterhazy, Saskatchewan. A surge of development by other companies resulted in twelve companies opening ten mines by 1970. In 1975 Potash Corporation of Saskatchewan was created by the government of Saskatchewan and shortly thereafter purchased several mines in Saskatchewan, including the Cory Mine (Cory), near Saskatoon. Cory has been in operation since 1968 and is located 12 km (7.5 mi) west of Saskatoon in central Saskatchewan. The mine is located in the center of the world’s largest potash deposit, 1,000 m (3300 feet) below surface. PCS has embarked on sizable capital expansions of more than CDN $7 billion. The Cory CDN $1.6 billion expansion was expected to raise operational capability of the site to 2.7 million tonnes per year, an increase of 1.9 million tonnes per year (MTPY). To achieve 2.7 MTPY of finished product and considering the 22.6% average production recovery1, 12 MTPY of ore must be hoisted to surface for processing. To deliver this rate required upgrading of the hoisting plant to 1500 tonne per hour (TPH). To accomplish this ambitious challenge, creative thinking, careful planning and a cohesive team was needed to work together to ensure the schedule, budget and most importantly the work was completed safely"

Jan 1, 2017

By Larry D. Patts

Section 317(e) of the Federal Coal Mine Health & Safety Act of 1969 directed the Secretary of the Interior to prepare standards under which all working places in a mine shall be illuminated by permissible&apos; lighting while persons are working in such places. Section 317(e) further provides that such Working places shall be illuminated within 18 months after such standards are promulgated. In accordance with this section of the Act, there was published in the Federal Register for December 31,1970 a notice of proposed ru1emaking which prescribed the illumination to be provided in the working p1 aces of underground coal mines. In 1ight of written comments, suggestions, and objections to this proposed ru1emaking, the proposed standards were with-drawn and reproposed in the Federal Register for Wednesday, October 27, 1971. In light of further comments, suggestions, and objections, a public hearing was held on April 4, 1974 and standards were again reproposed and published in the Federal Register for April 1, 1976. Promulgation of the final lighting standards took place on October 1, 1976 which means that the underground coal mining industry must comply with face illumination requirements by April 1, 1978. As mentioned previously, the first proposed ru1emaking for illumination of underground coal mines was published in the Federal Register on October 27, 1971. In early 1972, Consolidation Coal Company (Consol) and the United States Bureau of Mines agreed to a cooperative study of underground face illumination; Conso1 felt that expertise in this field would become increasingly important. Conso1&apos;s initial efforts in illumination were aimed at investigati.ng practical lighting systems for underground face equipment. We were concerned with installing unobtrusive lights which provided sufficient face illumination for safety, but at the same time were readily maintainable, electrically reliable, and physically sheltered from damage.

Jan 1, 1977

By Juan Andres Jarufe, Carlos Cifuentes

Numerical modelling has been extensively used to estimate plastic displacement in faults and rock mass surrounding mine excavations. Particularly, for fault slip seismicity, modelled plastic displacement is usually utilized to estimate seismic potential, however these methods assume that all the displacement caused during an specific simulated excavation step, is released as a single event at each step. This representation corresponds to a conservative estimation since rock mass displacement does not occur as one single event but as an exponential distribution of many small displacements and a few large ones. When displacement is expressed as seismic magnitude, this exponential distribution is addressed as the Gutenberg Richter (GR) relationship. This paper present a method to decompose the results from numerical models into a distribution of seismic events The reconstruction of the GR relationship from modelling displacement permits not only a more realistic approximation of the expected maximum magnitudes but also provides a tool to develop a probabilistic estimation of the maximum possible magnitudes.

Apr 26, 2022

By Michael Roach

"There’s a bit of confusion and misinformation in the tunneling and underground construction industry regarding contractor joint ventures. If it was all so clear I would not have been "asked" to write this article. For those of us on the contractors’ side of the fence, joint venture business relationships have been a part of life for a long time. But even for contractors, there can be a bit of confusion, and it normally begins on bid day.How often have you received bid results and wondered (out loud, usually), "Who the hell are the Lake Mead Intake Constructors"? or the "Regional Connector Constructors"? Once you get past the name, you see that there is nothing behind Oz’s curtain other than a couple of contractors bidding and building work in an attempt to make a profit. So take a few minutes to read the rest of this, and I’ll do my best to dispel a few mythWhat is a joint venture?In some form or fashion, joint ventures are as old as industry itself. In the United States, some of the earliest (and largest) joint ventures involved the railroads and oil companies in the late 1800s. In the United States’ underground construction industry, joint venture names such as, Healy-Greenfield, Green and Winston, Gates & Fox and Walsh, Shea-Ball, Kenny-Paschen and S&M-Traylor bring back memories of successful joint ventures that operated over many projects.There are nearly as many definitions for joint ventures (JVs) as there are attorneys. A short and simple definition is, "an association of two or more companies engaged in a singular, for-profit business enterprise without actual partnership or incorporation." That’s it. Two or more contractors decide to partner up for one or more projects. One benefit of the JV relationship is that the partners typically do not need to file incorporation documents, or articles of organization within the state where they intend to operate. So it is a relatively easy business arrangement to pull together. Other benefits of the JV arrangement include: • Oportunities to gain new expertise and capacity.• The ability for firms to move into new geographic markets.• JV agreements are not long-term commitments.• Sharing of financial support for a project.• Sharing of financial risks within a project. between the partners.• Can complicate dispute resolutions with the owner and between partners"

Jan 1, 2014

By W. E. Adkins

Ore and waste haulage usually constitute one of the major cost items in the operation of an open pit mine. This has always been true at the Chino mine, one of Kennecott Copper Corporation&apos;s western mines, located in southwestern New Mexico. The Chino Mines Division is an integrated operation. The open pit mine, dump leaching operation, and a precipitating plant are located at Santa Rita, New Mexico, and a power plant, mill, smelter, and fire refinery are located in Hurley, New Mexico, about nine miles south of the mine. Approximately 22,000 tons of ore are mined and processed each day; in addition, approximately 42,000 tons of waste are removed from the pit. Originally the pit was located along the side of Santa Rita Creek but it has deepened and expanded so that at the present time, it could better be described as a nearly circular pit about one to 1.5 miles in diameter. The bottom is at an elevation of 5750 feet, the west rim near the ore gathering yard is at an elevation of 6230 feet, and the highest bench on the east side of the pit is at an elevation of 6750 feet, giving a vertical mining depth of 1000 feet. The train waste dumps reach a maximum elevation of 6580 feet. At the present time, Chino uses electric locomotives, off-the- highway trucks, and a skip system for mine haulage.

Jan 1, 1963

By Edward R. Mueller

As an alternative to a million dollar, sensor based, computer dispatching system, Cyprus Pima Mining Company has installed a home¬made analog computer. The device is used in conjunction with a standard radio-dispatching system; it is inexpensive and very simple to build and operate. It enables the dispatcher to make near opti¬mum dispatching decisions and to identify all off schedule haulage trucks. Since installation, Cyprus Pima has experienced significant increases in pit productivity.

Jan 1, 1977

By S. F. Willey

The difficulties encountered in justifying replacements for over-age mobile mine equipment at the Gay Mine are common to most mining operations. More critical management reviews, resulting from ever rising capital expenses, increased such difficulties. These increased difficulties are especially apparent to those who are primarily operators and are not statisticians, mathematicians, or computer oriented. The Gay Mine is located 17 miles east of the town of Fort Hall in southern Idaho. This open pit operation annually supplies the J. R. Simplot Company and the FMC Corporation with 1.5 MM - 2.0 MM tons of phosphate ore. Stripping ratios average approximately three bank cubic yards of waste per ton of ore mined. The mobile mine equipment fleet consists of ten 24 yard scrapers; twenty-one 30-50 ton end dump trucks; four 32 42 yard shovels; fifteen dozers in the D-8 and D-9 classes; and three road patrols. Equipment replacement is based on the results of evaluations made on individual units. Authorization of capital expenditures for equipment replacement is dependent upon a favorable profitability index (true rate of return on investment) plus a favorable payout time. The system is based on cash flows and requires the determination of (1) the net cash outflows in the form of capital expenses plus (2) the expected cash inflows resulting from reduced expenses and depreciation tax credits. After adjustments for taxes and the time value of money, the net amounts are used to calculate the required indices. The basic elements for the system related to the replacement of mobile mine equipment include: (1) the determination of dollar penalties to be expected from the continued use of aging equipment; (2) the selection of equipment that will provide the greatest return on the investment dollars; (3) the quantification of the expected savings; and (4) the calculations of the prescribed indices. Inadequacies inherent in previous evaluating methods led to the search for more

Jan 1, 1972

By D. W. Mitchell

Never increase, reduce, or remove ventilation without unquestionable, compelling reasons. This rule applies every day as well as when fighting fire, attempting to find survivors, or recovering the mine. Right now in some mine a miner is adjusting a regulator in one place to get more air into another. The chances are the desired result will be achieved. Has the miner done a good thing? NO! Solving the immediate problem reinforces the belief he or she can do it again. That "again" might be, to give but one of many examples, the explosion in Greenwich No. 1 Mine. Robbing the proverbial Peter can be the prelude to disaster in a mine. Reducing the airflow is a common practice when fighting a mine fire. In fact, some people have been taught to do that at MSHA&apos;s academy. People who believe that to be proper must be people who never experienced a fire. To fight most fires you have to scramble to get enough air. An explosion demands totally different "thinking". If people are soon to go into the mine, DON&apos;T adjust the ventilation. If teams are in the mine, each affected area must be explored or sealed before making any ventilation change.

Jan 1, 1993

By V. P. Petrov

There is a great number of asbestos deposits in the USSR. Some of them are quite large ones. The main deposits of chrysotile asbestos can be seen in the sketch map (fig. 1). Bazhenovo deposit situated in Sverdlovsk region is the largest of them. It is on mining operations since year 1889. Systematical studies of asbestos deposits were begun from the first years of Soviet Government and connected with the names of K.E. Tarasov, P. M. Tatarinov, B. J. Merenkov, N.D. Sobolev, F. V. Syromjatnikov, V. N. Lodochnikov, V.P. Petrov, V.P. Eremeev, L.A. Sokolova, V. R. Artemov, V.F. Dybkov and many others. This investigations served the USSR to be one of the most important producers of asbestos in the world. In the limits of the report a characteristic of the main asbestos deposits of the USSR may be done in general outline, from the point of view of their genesis, and in order to show their common features and differences that must be taken into account for forecasting, exploration, and mining of asbestos. All chrysotile asbestos deposits of the USSR were formed of magnesium rock as a result of comparatively lower-grade temperature hydrothermal alteration and free crystallization of fibrous chrysotile in open cavities that served as catalyst As a matter of fact an important mineral for industry is not chrysotile in general, a widely distributed rock-forming mineral, but a long fibrous and mainly vein variety of chrysotile. It must serve for geologists as an important prospecting criteria. A geologist should search for the mineral not only places of possible chrysotile generations, but places where asbestos may be crystallized in its long fibers that could be fiberized well. The Soviet geologists distinguish two groups of chrysotile-asbestos deposits: (1) related to ultrabasic rocks and (2) related to sedimentary dolomite terranes, taking into account what the maternal rock was affected by hydrothermal solutions. The largest deposits related to ultrabasite. The largest deposits in dolomite are small ones but their asbestos contains almost no iron. So much asbestos is a better filler into insulating plastics. There is a few types of asbestos deposits in the limits of the ultrabasic group. The largest of them is a Bazhenovo type which the deposits of Urals are attributed, and Bazhenovo deposits is the first instance. The Bazhenovo type is characterized by the most complete zonality in disposition of various kinds of asbestization. This type of deposits is practically the only one to be in working now in the USSR.

Jan 1, 1978

By M. R. Shahan, W. R. Reed, P. Zhang, J. P. Rider

Roof bolter operators may be exposed to high respirable dust concentrations on continuous miner sections with blowing face ventilation when bolting is performed downwind of the continuous miner. One solution to reduce the high respirable dust concentrations is to use a canopy air curtain (CAC) to deliver clean air from a filtered blower fan directly to the bolter operators under the canopies. The influence of CAC installation in the airflow and dust dispersion around the location of the roof bolter operator can be evaluated by using computational fluid dynamics (CFD). This study, performed by the National Institute for Occupational Safety and Health (NIOSH), considers two scenarios: (1) a roof bolting machine in the center of the entry for installation of the fifth row of bolts from the face, and (2) a roof bolting machine positioned close to the face for the installation of the last row of bolts. In both scenarios, the bolting machine is placed in an environment which contains 6.0 mg/m3 of respirable dust and is ventilated by a blowing curtain with 3,000 cfm (1.42 m3/s) of air. This environment is used to simulate the roof bolter machine operating downstream of a continuous mining machine. Two operation positions are simulated at the same bolting location: dual drill heads in the inward position for two inside bolts; and dual drill heads in the outward position for two outside bolts. The influence of the CAC on airflows and dust dispersion is evaluated with the CAC operating at 250 cfm (0.12 m3/s). INTRODUCTION Roof bolting machines are used to drill bolt holes into the roof of an underground mine at specified patterns and to install roof bolts to secure the roof from falling. At times, the roof bolting machine is also used to install rib bolts; however, only the roof bolting operation was considered in this simulation. Though the roof bolting operation drastically improved the safety of miners from roof falls, some roof bolter operators are experiencing health problems associated with respirable dust. Based on recent statistics [1], following a low point in the late 1990s where the prevalence of coal workers’ pneumoconiosis (CWP) in miners with 25 years or more tenure was approximately 5%, CWP now exceeds 10% for miners with 25 years or more of tenure. In central Appalachia (Kentucky, Virginia, West Virginia), 20.6% of long-tenured miners have CWP and 4.5% have progressive massive fibrosis, the most severe form of CWP. CWP, also known as black lung, is an occupational respiratory disease that has no cure and can ultimately be fatal. The only way to prevent the incidence of black lung is to eliminate miners’ exposure to respirable coal mine dust. Current U.S. regulations require that the overall respirable dust concentration to be 1.5 mg/m3 or less for a full-shift sampling if respirable quartz is not present; otherwise, the standard is 10 divided by the percent quartz (not to exceed the 1.5 mg/m3 respirable dust standard) [2, 3].

Jan 1, 2019