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By J. G. Clevenger

Introduction Computer-based haul truck dispatching has been operational at Tyrone since May, 1980. The system, known as Dispatch, has accomplished significant productivity gains compared with the former method of manual radio dispatching. These gains translate into a reduction in equipment requirements or an increase in production. An operational mine may choose to increase production or reduce the number of equipment shifts operated, resulting in substantial operating savings. Dispatch Description The core of Dispatch is a VAX 11/780 computer. Originally purchased with 512 kbytes of memory, the system is now upgraded to 3.75 Mbytes to accommodate users from all areas of the operation. Dispatch alone requires 200 kbytes of locked-in memory, with an additional 300 kbytes of memory available on a top priority basis. The computer, located in the Tyrone Branch general office building, is connected by under-ground cables to the dispatch tower, 3.2 km (2 miles) from the general office. Four two-paired lines are used to drive two cathode ray tubes, a printer, and the Dispatch master panel. One CRT continuously displays pertinent pit information for the dispatcher's use. Routinely displayed transactions include truck assignments, changes in shovel material type, truck and shovel breakdowns and delays, and changes in truck, shovel, or dump status. Er¬ror detection capabilities include late arrival and improper input from the field panels. The second display terminal is used for utility reporting and input functions. Status reports are displayed for availability, position, and material type being handled by any given truck or shovel. Fuel reports are available indicating the time last fueled, quantity of fuel delivered, and the time spent fueling. A truck factor utility reports relative performance to other trucks and a shovel utility reports efficiencies in tons per shovel shift. Instantaneous shift reports and mine status reports are available. Inputs such as lineups, operator changes, new travel times, shovel bias, tie down times, and lunch times are all handled with the utility display terminal. The printer records a hard copy of operating truck locations and production from the beginning of each shift. If there is a system malfunction, the printer copy is readily available to the foreman for radio dispatching. The Dispatch master panel is a microprocessor-based panel that allows the dispatcher to override any computer-assigned transaction. Classification of "bad order" conditions and acceptance of certain field messages is entered through the master panel. This panel communicates with all truck and shovel field panels via UH radio frequency through a repeater tower located on a high ridge contiguous to the mining area. Each truck and shovel field panel unit transmits a coded signal that is identified by the computer. A typical communication sequence between field panels and the computer for one complete truck cycle is: • The truck panel LED display indicates the shovel number to which it is assigned. The truck arrives at its assigned shovel and the operator presses the "Arrive" button on the truck panel. As the computer receives the "Arrive" message, it concludes the return (empty) travel time and begins accumulating idle time before loading. • When the first dipper of muck is placed into the bed of the truck, the operator presses the "Load" button. The computer concludes the truck idle time accumulation and begins the loading time calculation. • When the truck is loaded, the shovel operator presses the "OK"

Jan 9, 1983

By Dennis Z. Mraz, Maurice B. Dusseault

INTRODUCTION In 1982 and 1983, as part of the mine design program for a New Brunswick potash mine, series of laboratory tests were conducted on potash and salt samples obtained during the mine exploration program. Concurrently, two rock mechanics monitoring programs were initiated in-situ the short term test at site A, and the long term test at site B (figures 1 and 2). SHORT TERM TEST (FIGURE 1) The purpose of this test was to obtain a quick confirmation of laboratory data and comparison of the rock response to excavation with the data available from Saskatchewan potash mines. Instrumentation in the main drift was installed on Day 4 after excavation, and the pressure cells were pressurized on Day 33. The packer type cells in the east crosscut were pressurized on Day 25. Excavation of the west cross-cut was begun on Day 71. This "mine by" test has provided important information required for calibration of laboratory tests, as will be shown later. The test data recorded are shown in a series of diagrams: a. initial room closures, main drift-figure 3; b. convergence graphs, main drift-figure 4; c. stresses recorded, main drift-figures 5 to 8 and, d. stresses recorded, east crosscut - figure15 Three types of instruments were used in this test. In the main drift. Saskatchewan Research Council water type hydraulic inclusions. encapsulated in potash, were used to measure lateral and vertical stresses. In the east crosscut a home-made packer type hydraulic inclusion was used to record radial stresses. The convergence stations consisted of end-grouted bars of reinforcing steel. LONG TERM TEST (FIGURE 2) The purpose of the long term test was to study the steady-state behaviour of a single opening in potash. Site B was placed in a remote area of the mine, well away from any anticipated early mining activities. To assure homogeneity of rock mass, the site was placed within a 30 meter thick vertical limb of a syncline-shaped ore body contained within a 360 metre thick vertical section of salt rocks. It was excavated mechanically as close as possible to the designed square cross-section for the entire length of the room. The square cross- section was chosen in order to achieve a comparable geometry in vertical and horizontal directions; that is, to eliminate any undesired effects of geometry interaction with lithostratigraphy which might have generated non-homogeneous behaviour. The site was extensively instrumented. As in site A, the convergence stations consisted of end-grouted bars of reinforcing steel. However, in order to avoid any measurement inaccuracies which would be associated with the brittle yield zone in the opening walls, bars were placed to the same depth in opposite sides of the opening. This allowed the convergence measurement to be obtained as the absolute convergence between two points, rather than the convergence of a single point with respect to a potentially unstable reference. For stress measurements the Saskatchewan Research Council cells were placed vertically, laterally and radially. Unfortunately, due to cost considerations, it was decided to use a compacted sand, rather than grout, to encapsulate the cells in the boreholes. Due to the extremely small strain rates, only radial ly placed cells could be adequately pressurized. The rest of the cells did not perform adequately, and are currently being replaced.

Jan 1, 1986

By David R. Delling

Introduction In 1967, Tenneco acquired Kern County Land Co. The primary reason for the acquisition was to obtain Kern's agricultural, manufacturing, and oil interests. Included in the Kern portfolio were the core of Tenneco's present sodium leases in Sweetwater County, WY. This lease base has been ex¬panded and now consists of 62 km2 (24 sq miles). Beneath this property lies more than 500 Mt (550 million st) of economic trona ore. Following the Kern acquisition, Tenneco continued exploration and feasibility studies concerning the trona leases. The effort was accelerated in 1976. Tenneco was attracted to trona because of the favorable lease position it held, the growth rate of the Wyoming industry during the 1960s and 1970s, and the prospects for a long-term, stable business enterprise. The reserve picture was certainly attractive to a company long accustomed to the never-ending struggle to replace oil and gas reserves. In 1978, Wyoming was a boom area that had experienced rapid growth for a number of years. Unemployment was very low and vacant housing nonexistent. The prospects for soda ash were somewhat different in the 1970s than now. But it was recognized even then that a new producer would have a challenging task to compete in this business. Tenneco knew the capital burden required to build a grassroots facility during inflationary time, would have to be offset elsewhere Since labor and energy are the dominant direct costs in mining trona and producing soda ash these factors have always received a great deal of attention. Contracting and permits Environmental monitoring on the site began in late 1977. With the already mentioned considerations in mind, engineering began by mid-1978. Brown & Root Inc. was selected as the general contractor for the project. S. A. Scott & Co. was the engineering subcontractor for the mine shafts, hoists, ventilation system, and miscellaneous mine facilities. Frontier-Kemper was awarded the contract for sinking the two mine shafts. Acquisition of numerous federal, state, and local permits occupied all of 1978 and 1979. The Wyoming Industrial Siting Permit, which addresses socioeconomic issues, drew the most public interest. Siting issues were all satisfactorily resolved by negotiating reasonable stipulations with various local governmental units, school districts, and the like. Site location The Green River project location is not truly remote. It is located within 5 km (3 miles) of Interstate 80 and only about 32 km (20 miles) west of the city of Green River. Nevertheless, a substantial investment had to be made in communication and utility links to the site. These included: • A 14-km (8.5-mile) rail spur connecting with the Union Pacific Railroad at Bryan, WY. This spur has four bridge structures including one crossing Interstate 80. Most soda ash is shipped by rail, and coal is received over this line as well. • A paved 5-km (3-mile) roadway connecting to Interstate 80. • A powerline extension of about 27 km (16.5 miles), along with two new substations. • A 24-km (15-mile) waterline connecting the plant to a pump station on the Green River. The site is located on gently sloping terrain at an elevation of about 2.1 km (6800 ft). Ground conditions are good for construction. Weathered bedrock lies near the surface in most places. Geology The general geology of the Green River Basin has been well described previously. Tenneco's ore is in several superimposed evaporate beds. But the bulk of it is present in Bed 17, located about 457 m

Jan 10, 1985

By S. K. Kawatra, H. Haselhuhn

"In many mineral processing applications it is necessary to know the isoelectric points of the various minerals in the ore to properly perform flotation and other mineral separations. However, current zeta potential analysis technology used to determine the isoelectric point of a particle suspension cannot deliver the isoelectric point of each individual mineral, just an average of the mineral system. Traditionally, individual isoelectric points have been gathered by testing pure, synthetic minerals. These values can vary widely depending on the mineral synthesis procedure and the water quality used during the isoelectric point analysis. A more robust approach for determining the actual isoelectric point of a particular mineral in a mixture of minerals is needed. This paper details a method for determining the isoelectric point of an iron oxide mineral in a siliceous iron ore. The method uses a laser scattering particle size analyzer and a pH electrode to determine the pH at which the liberated mineral particles begin to flocculate as pH is decreased. The pH at which the average particle size rises dramatically is the isoelectric point of the mineral with the highest isoelectric point in the ore. This measurement technique was used on natural hematite, goethite and siderite ores as well as a synthetic mixture of pure silica and pure hematite. The results for synthetic hematite mixtures were comparable to literature values.IntroductionRecently there has been an increased interest in the processing of low-grade iron ores (Bolen, 2014; Carlson and Kawatra, 2008, 2011, 2013; Halt et al., 2014; Halt and Kawatra, 2014; Haselhuhn 2012, 2013; Haselhuhn et al., 2012a, 2012b; Kawatra and Halt, 2011; Liu et al., 2014; Manouchehri, 2014; Sandvik and Larsen, 2014; Semberg et al., 2014). The zeta potential of iron ore minerals is a very important and often overlooked aspect of iron ore mineral processing. Zeta potential is the charge at the shear plane of a surface in an aqueous environment (Carlson and Kawatra, 2013). Many ores consist of a variety of minerals differing in zeta potential. This property of a particle surface in water is responsible for either keeping the particles dispersed from each other or flocculated to each other. This is important during many operations such as flotation, desliming and thickening. Zeta potential is primarily a function of the pH of the water and the mineralogy of the particles, though certain ions within the water can also have a significant impact (Haselhuhn, 2012). Consider the following situation: silica (SiO2) particles and hematite (Fe2O3) particles are suspended in an aqueous solution. These two types of minerals have drastically different zeta potentials throughout the pH spectrum (Esumi et al., 1988). In alkaline conditions, both types of particles exhibit negative zeta potentials and will tend to disperse from each other. In acidic conditions (between pH of 3 and 6), silica has a negative zeta potential while hematite has a positive zeta potential, so these surfaces will tend to be attractive to each other and the particles will flocculate and settle. This phenomenon is visually described in Fig. 1. The pH at which the two minerals begin to attract to each other, or destabilize, is near hematite’s isoelectric point. The isoelectric point is defined as the pH at which the surface has a net zero zeta potential (Carlson and Kawatra, 2013)."

Jan 1, 2015

By T. M. Klemetti, T. J. Batchler, T. J. Matthews

How representative are partial-panel screen tests of the in situ response of screen to roof and rib deterioration and displacements? Can a more representative procedure that tests up to two full screen panels improve the utilization and effectiveness of screen as a roof and rib support? This paper details a new full-panel screen test procedure and the initial results. The full-panel test frame was designed to test a 2.4 by 3.6-meter screened area with bolt spacing as small as 0.3-meter increments. The initial design included up to 6 screen pull points, but additional locations can be added to produce a more uniform loading. The test frame was developed to capitalize on the capabilities of the Mine Roof Simulator (MRS) at the Pittsburgh Mining Research Division (PMRD) of the National Institute for Occupational Safety and Health (NIOSH). The initial tests included a baseline test of a partial panel (1.22 by 1.22-m screen section) using the original 1.52 by 1.52-meter test frame, a 1.22 by 1.22-meter screen section test in the new full-panel test frame, and two tests of 2.4 by 3.6-meter screened area in the full-panel test frame. The 1.22 by 1.22-meter screen section tests in the full-panel test frame produced a peak load 6% higher and a yield load 30% lower than the partial-panel (original) test frame. The multiple pull points of the full-panel frame produced variable yield, peak, and intermediate loads, but on average were lower than the single pull-point tests. INTRODUCTION Roof and rib surface support has become a major component of the stability and safety of underground coal openings in the past 15 years. The use of screen, either metal or synthetic, has proven to decrease the injuries of underground coal miners. By 2005, there were 466 injuries in the United States caused by roof skin failures, with numerous mines not employing roof screen. In contrast, in 2015 there were 175 injuries caused by roof skin failures, with at least some of the reduction due to the increased use of roof screen. A continued effort to increase the utilization of screen in areas of underground coal mines where workers are traveling and working can reduce these injuries even further. To aid in this effort, the testing of various screen types under various loading and attachment conditions is required. Past testing efforts have primarily consisted of partial-panel testing in a laboratory setting. Laboratory testing of roof screen, both metal and composite, has been conducted in several ways by researchers, manufacturers, and other entities. The primary focus on the testing has been evaluating weld integrity of 1.22 by 1.22-meter screen section tests of various screen types utilized in mining (Dolinar, 2006; Dolinar, 2009; Tannant, 2001). There have also been several efforts at numerical simulation of the screen and screen/rock interactions (Murali et al., 2006; Shan et al., 2014). A few full-panel tests were performed with one pull or loading point in the center of the screen panel (Shan et al., 2014). As most previous efforts have mentioned, the only true way to evaluate the capabilities of screen is to conduct full-panel testing. Single wire strand, weld, and partial-panel (1.22 x 1.22-m) tests as well as numerical modeling only provide insight into the actual response of full-panel screen. A new test frame and procedure have been developed at PMRD to conduct various loading, anchoring, and material testing.

Jan 1, 2019

By R. J. Matetic, J. Noll, C. Zhou, J. DuCarme, M. Reyes, J. Srednicki

"In April 2016, MSHA began requiring the use of continuous personal dust monitors (cPDMs) to monitor and measure respirable mine dust exposures to underground coal miners. After the cPDM’s implementation, mine operators discovered that it interfered with proximity detection systems (PDSs), thus exposing miners to potential striking and pinning hazards from continuous mining machines. NIOSH was sought out by MSHA and mining industry stakeholders to determine how the cPDM and PDS interact with each other. Accordingly, NIOSH investigated existing standards, developed test protocols, designed experiments, and conducted lab evaluations. Some interferences were observed to be caused by electromagnetic interference (EMI) from the cPDM. Results showed that there was no significant interference when the cPDM and the miner-wearable component of the PDS were separated by distances of 6 inches or greater. In this study, the cPDM and PAD needed to be at least 6 inches apart in order for them to be used simultaneously and reduce interference potential. INTRODUCTION Underground coal miners are exposed to a variety of hazards on a daily basis such as coal dust exposure, high noise levels, roof and rib falls, the potential for fires and explosions, and operating and working with heavy machinery. One of the hazardous jobs is operating or working nearby a continuous mining machine (CMM). According to Mine Safety and Health Administration (MSHA) statistics, 40 miners have been fatally struck or pinned by a CMM since 1984. In an effort to prevent future striking and pinning fatalities from occurring, proximity detection systems have been developed and are required on all operating CMMs in underground coal mines, with the exception of full-face CMMs, by 2018 (1,2). Proximity detection systems (PDSs) are designed to alarm miners and stop machine motion in order to protect miners from being struck, pinned or crushed by CMMs(3). Currently, MSHA-approved PDSs, installed on CMMs, are based on the principle of magnetic flux density (B-field) (4,5). The system generates a magnetic field around a CMM and determines the relative distance a miner is from the CMM based on a detected magnetic flux density. A PDS typically consists of multiple magnetic field generators mounted at different places around the CMM, and personal alarm devices (PADs), which are worn by the miners to detect the magnetic flux density. When a miner, wearing a PAD, gets closer to the machine, the PAD detects a stronger magnetic field from the generators and detects a weaker magnetic field when a miner moves away from the machine. The magnetic field generators installed on CMMs typically produce modulated magnetic wave signals at frequencies between 10 kHz and 120 kHz. The magnetic fields are measured by three small magnetic coil antennas mounted on orthogonal axes inside of the PAD worn by the miner. According to Faraday’s Law, the changing magnetic field induces a voltage in each coil antenna, and the voltage values are transmitted from the PAD back to a controller on the CMM wirelessly, typically at a frequency between 400 MHz and 2.5 GHz. These values are then used to determine the distance between the miner and the generators of the machine. Typically this information is used to determine when a miner wearing a PAD is in a warning zone or stop zone which would trigger different alarms and actions such as slowing the machine down or stopping it."

Jan 1, 2017

By Isobel Yeo, Luca Cocchi, Sven Petersen, Nico Augustin, Sean Plunkett, Ingo Klaucke

Exploration for seafloor massive sulfides is currently mainly focused on investigation for active hydrothermal vent fields. However, most of the resource potential is thought to be related to inactive, large deposits that had enough time to accumulate sufficient tonnage. At the same time, ageing of these deposits will ultimately result also in burial acted by either sediments and/or volcanic deposits. We currently lack the technologies to explore for these deposits in a time- and cost-efficient manner. In the fall of 2012 we used a number of AUV-based technologies to explore and image a sulfide occurrence at the Palinuro Volcanic Complex in the Tyrrhenian Sea (Mediterranean Sea; Fig. 1) where sampling and drilling have documented a buried polymetallic barite-pyrite dominated site (Petersen et al., 2014). Sulfide mineralization is here covered by volcaniclastic and pelagic sediments and only red sediment staining and the presence of few tubeworm bushes indicated recent hydrothermal activity - black smoker venting is absent. We used AUVbased high-resolution bathymetry, sidescan sonar imaging, subbottom profiling, as well as magnetics to describe, in a quantitative way, the previously known sulfide-bearing areas as well as to successfully explore for new hydrothermal sites including large chimney structures (Figs. 2). Additionally chemical sensors (CTD, Eh, turbidity) investigated the water column for signs of hydrothermal activity in areas not previously investigated, but at the end, only weak anomalies were found in the southern part of the summit.

Sep 1, 2014

By M. Schafer

The 6-m- (20-ft-) diameter cast-in-place tunnel lining for Phase 3 of the Mill Creek Tunnel (MCT3) has been completed. This article summarizes aspects of liner design and ad-dresses the construction approach in placing the liner in areas of tunnel that experienced crown overbreak. Tunnel lining production rates, installation procedures and form buoyancies experienced during concrete placement will also be presented. The Mill Creek Phase 3 tunnel (MCT-3) is expected to be completed this year. It is one of the largest tunneling projects undertaken by the Northeast Ohio Regional Sewer District (NEORSD) to date. The tunnel horizon is situated within the Devonian Chagrin Shale rock formation at an average depth of 85 m (280 ft). The tunnel was excavated using a two-pass method. A full face, fully shielded, Robbins 7.2-m- (23.8-ft-) diameter tunnel boring machine (TBM) was used to excavate approximately 4,500 m (15,000 ft) of tunnel and facilitate installation of initial supports (first pass). The final lining (second pass) consists of 305-mm- (12-in.-) thick cast-in-place reinforced concrete and integral low flow channel. Seven shafts were constructed on the project. The excavation of the tunnel commenced at Shaft 14 and proceeded down grade to Shaft 9, the terminus shaft. Because of the presence of overbreak in the tunnel crown, measures were required to support the form and counteract the effects of buoyancy during concrete placement.

Jan 1, 2008

By Verya Nasri

The proposed Second Avenue Subway Project in New York City is approximately 8.5 miles long, up to a depth of 120 feet beneath grade and runs the length of the eastern side of the island of Manhattan from Harlem in the north to the financial district in the south. It will reduce overcrowding and delays on the existing parallel line, improving travel for both city and suburban commuters, and upgrading access to mass transit for residents of the East Side of Manhattan. This subway line will have 16 new stations providing transfers to the existing Metro-North commuter rail service and to New York City Transit subway lines and bus services. The first phase of construction includes two TBM rock tunnels 2.4 miles long and two new large and shallow rock cavern stations with clear span openings ranging from 70 to 100 feet and length in excess of 1000 feet with 615 feet center platform, coupled with low rock cover ranging from 20 feet to 40 feet. The design methodology for these large and shallow rock caverns consists of using well known empirical methods for initial estimate of support requirements and then two and three dimensional continuum and discontinuum analyses for verification of the selected excavation sequence and support system. This paper describes the analysis and design methodologies used for the two large and shallow rock caverns with insight into how to overcome the design complexities and construction challenges of the urban underground.

Jan 1, 2008

By Joseph J. Zaluski, J. Steven Gardner, Marco M. Rajkovich

The telephone rings at 3:20 a.m. An exasperated voice on the other end says, “There has been a mine explosion at Mine No. 5 and we have not heard from three units of men.” Twenty-three minutes later, the phone rings again and another panicked voice says, “There is something wrong at the slurry impoundment, the pool level has just dropped five feet and we are trying to figure out what is wrong” Thirty minutes later, the mobile phone rings. An excited security guard says, “We have reports of thick black water running down the right fork and left fork of Straight Creek and both seem to be rising in the banks.” If this is not enough excitement, the mobile phone rings again. A stressed company official says, “I just heard a report that black slurry is running out of the portals of XYZ Co. No. 2 Mine, which is adjacent to our No. 5 Mine.” Is this all some kind of dream or nightmare? More importantly, though, if this happens, what do you do? Most people think: “It can’t happen to me.” But mine explosions do happen, slurry impoundments can rupture and creeks can flood with black water causing damage to others, including competitors. A mine’s response must be quick, cool-headed, calculated and effective. This article examines the issues, concerns and ramifications of this hypothetical disaster. The authors then present a systematic crisis-management response.

Jan 1, 2004

By Douglas D. Campbell, Alan S. Imrie

INTRODUCTION Mica Dam is located on the Columbia River just downstream of the "big bend" of the Columbia, where it leaves the Rocky Mountain Trench and begins to flow southwards, approximately 85 miles north of Revelstoke, B.C. (Fig.l). The dam was built as one of three Columbia River Treaty dams in British Columbia which regulate the flow of the river and provide flood control downstream in Washington and Oregon. In addition, the three Canadian dams make possible greatly increased power generation at existing power plants in the United States and further development along the Columbia and its tributaries in British Columbia. History Mica Dam, constructed during the period 1967 to 1973, is a zoned earthfill embankment 650 feet high, which has 12,000,000 acre feet of live storage and an available head of 550 feet. Power intakes which would serve either a six-unit surface or underground plant were incorporated into the design and construction of the right abutment of the dam (Fig. 2). Prior to 1971 alternative 2500 MW capacity surface and underground powerplant arrangements were studied. Cost estimates for the two alternatives were the same within the limits of estimating procedures. However, it was believed there was a certain continuing risk in locating a surface plant at the toe of a steep canyon wall. Therefore, in 1971 a tentative decision was made to proceed with an underground plant, provided that additional exploration confirmed assumptions of

Jan 1, 1976

By Scott G. Britton

Microcomputeranalysis has long been a technique for engineering professionals to improve the theoretical and practical design of continuous mining systems. Such work has been typically composed of modeling observed events and then introducing a variety of known elements with predicted results to compare against the results of the model. Programs written in BASIC, Cobol, Fortran or other language dominated such efforts. These programs have taken years to develop and prove out. This article explores a new management tool used daily at Tanoma Mining Co. (TMC) to track and improve continuous mining systems through the use of microcomputers and Lotus l-2-3 software. Data provided by section foreman, shift foreman, and maintenance personnel are analyzed against predetermined goals. In this way, new production targets are developed, weaknesses in the current mining systems are identified, and the results of changes are measured. The current results of this system on the productive output of TMC systems is on the order of 10% improvement. Thin seam coal mining is an industry within the broader industry of coal mining. Thin seam mining faces the same cost pressures and engineering problems as its higher-seam competitors without the advantages of moving volume per foot of extracted material to stay at an economic parity with others requires additional creativity within the management teams to extract higher productivity gains from the existing workforce and mining systems. TMC operates a deep mine in Indiana bounty, PA, about one hour northeast of Pittsburgh.

Jan 1, 1991

This paper evaluates the effect of longwall mining on a cluster of unconventional shale gas wells located in a district barrier pillar between a current longwall panel and the future panel north of the barrier pillar. A cluster of 19 unconventional shale gas wells have been drilled and 5 additional gas wells have been proposed in the barrier pillar. The closest gas well is 25.6 m (84 ft) to the current panel tailgate and 26.2 m (86 ft) to the tailgate of the future panel north of the barrier pillar. The rib-to-rib width of the barrier pillar is 57.9 m (190 ft). The distance between the two gas wells across the barrier pillar is 6.1 m (20 ft). The pillars in the current tailgate are on 22.9-m (75-ft) and 32.0-m (105-ft) centers, and the pillars in the tailgate of the panel north of the barrier will be on 22.9-m (75- ft) and 36.6-m (120-ft) centers. The two longwall panels adjacent to the barrier pillar are 475.5-m (1,560-ft) wide. The overburden depth at the shale gas well site is 321.6 m (1,055 ft). The wells have been installed in the barrier pillar, and the current district panel and subsequently the panel north of the barrier will be developed and retreated in the near future. The average min

Jan 7, 2020

By Mina M. Shinouda

Brightwater West (BT4) represents the state of the art in utility tunneling, namely a long relatively small diameter, soft ground tunnel, with no intermediate shafts, under significant groundwater pressures, requiring very precise survey control in order to hit a small exit window. The tunnel is over 6.4 km (4 miles) in length, is expected to encounter active earth pressures of over 5 bars in glacial geology, and a planned hole through into a shaft eye constructed at 45.7 meter (150 feet) below the water table. Despite the fact that all of these challenges have been previously overcome in larger diameter tunnels, the technical solutions are much more challenging in a smaller diameter due to the lack of space available for implementing the equipment and techniques required. In addition to the technical challenges, the project is faced with many of the constraints designed to minimize the impact of these types of projects on the neighboring community and environment. The work shaft for BT4 is within 15.2 m (50 feet) of Puget Sound and is on a site that was previously occupied by petroleum storage tanks, and all of the tunnel muck has to be removed from the site by barge as trucking is not allowed by contract. The paper addresses the work performed during the preparation for tunneling stage of this project both in assembling the TBM and in preparing the site for its arrival.

By C. D. Litton

Laboratory experiments were conducted to determine the responses of a prototype smoke detector and a commercially available photoelectric smoke detector to smoke particles generated from various combustion sources. The prototype smoke detector combines optical scattering measurements with ionization chamber measurements in order to reduce/eliminate nuisance alarms due to the presence of airborne dusts or diesel exhaust particles. The commercially available smoke detector is designed for use in harsh environments where airborne dust represents a major problem due both to nuisance alarms and detector contamination. In the experiments, the responses of the two detectors were measured when exposed to smoke particles from the exhaust of a diesel engine and from a variety of fire sources, including wood, coal, styrene butadiene rubber, and No. 2 diesel fuel. For the solid fuels, data were obtained for both smoldering and flaming combustion. This report describes the experiments, their results, and the use of these results as they apply to early-warning fire sensors capable of the rapid and reliable detection of fires in atmospheres that may or may not be contaminated by either airborne dust or the products produced from diesel engines. Subsequent to these laboratory experiments, six smoke detectors were installed in an underground limestone mine for further testing and evaluation. DISCLAIMER: The findings and conclusions in this report have not been formally disseminated by the National Institute for Occupational Safety and Health and should not be construed to represent any agency determination or policy.

Jan 1, 2009

By Miguel Reyes, Bruce Whisner, Joseph Bickson, Lincan Yan, David Yantek, Justin Srednicki, Nicholas Damiano, Eric Bauer

Built-in-place (BIP) refuge alternatives (RAs) are designed to provide a secure space for miners who cannot escape during a mine emergency. Heat and humidity buildup within RAs may expose miners to physiological hazards such as heat stress. To minimize the risk of heat stress, Title 30 Code of Federal Regulations (CFR), or 30 CFR, mandates a maximum allowable apparent temperature (AT) for an occupied RA of 35 °C (95 °F) (MSHA 2008 [1]). The National Institute for Occupational Safety and Health (NIOSH) has conducted extensive research on the thermal environment of occupied RAs intended for use in underground coal mines. NIOSH research has demonstrated that a fully occupied BIP RA can exceed the AT limit by > 5.6 °C (10 °F) in mines with elevated mine strata and air temperatures (Bissert et al. 2017 [2]). In this circumstance, an RA cooling system could provide a solution. This paper provides an overview of test methodology and findings as well as guidance on improving the performance of a cryogenic air system prototype by optimizing the flow rate, increasing the tank storage capacity, and improving the efficiency of the heat exchanger of the cryogenic system. This may enable BIP RAs to meet the 35 °C (95 °F) AT limit in mines with elevated temperatures. The information in this paper is useful for RA manufacturers and mines that may choose to implement a cryogenic air system as a heat mitigation strategy

By Miguel Reyes, Bruce Whisner, Joseph Bickson, Lincan Yan, David Yantek, Justin Srednicki, Nicholas Damiano, Eric Bauer

Built-in-place (BIP) refuge alternatives (RAs) are designed to provide a secure space for miners who cannot escape during a mine emergency. Heat and humidity buildup within RAs may expose miners to physiological hazards such as heat stress. To minimize the risk of heat stress, Title 30 Code of Federal Regulations (CFR), or 30 CFR, mandates a maximum allowable apparent temperature (AT) for an occupied RA of 35 °C (95 °F) (MSHA 2008 [1]). The National Institute for Occupational Safety and Health (NIOSH) has conducted extensive research on the thermal environment of occupied RAs intended for use in underground coal mines. NIOSH research has demonstrated that a fully occupied BIP RA can exceed the AT limit by > 5.6 °C (10 °F) in mines with elevated mine strata and air temperatures (Bissert et al. 2017 [2]). In this circumstance, an RA cooling system could provide a solution. This paper provides an overview of test methodology and findings as well as guidance on improving the performance of a cryogenic air system prototype by optimizing the flow rate, increasing the tank storage capacity, and improving the efficiency of the heat exchanger of the cryogenic system. This may enable BIP RAs to meet the 35 °C (95 °F) AT limit in mines with elevated temperatures. The information in this paper is useful for RA manufacturers and mines that may choose to implement a cryogenic air system as a heat mitigation strategy

By C. Sun

The Leeville underground mine has been in operation since 2006 and currently produces 3,500 tons of ore per day with an average gold grade of 0.34 opt. The mine has a large horizontal extent of 2,500 feet by 500 feet and a vertical extent of 20 to 200 feet. Based on the characteristics of the ore bodies, a long-hole stope and cut-and-fill mining methods are used at the deposits which are divided into several mining zones. Cemented rock fill (CRF) is used as the backfill in the open stope. CRF design includes the aggregate design, the ratio of cement and fly ash design, and strengths for the long-hole stope and cut-and-fill. In the aggregate design, the particle size distribution is one of the most influential aggregate characteristics in determining how it will perform as a basic backfill material. Gradation helps determine durability, porosity, cement and water requirements, strength, and shrinkage. To cut the backfill cost, fly ash usually is added into the binder to replace some part of cement. However, adding fly ash can affect the strength of the backfill. The reasonable percentage of fly ash is also important for making a good and low cost backfill. Based on the different mining method, the different backfill with different strength is designed. A systematic quality control procedure for aggregate and backfill is followed. After four yeas application, CRF system in Leeville achieved very good results.

Jan 1, 2011

By Jacek Paraszczak

Numerous problems encountered in the mining of narrow vein deposits challenge the productivity, competitiveness and profitability of many mines. Mechanization of mining operations is one of the most important means to ensure the healthy survival of this mining sector. Vein deposits account for nearly all the world's primary silver production, as well as a significant portion of gold, tin, copper lead and zinc. In North America, narrow vein mining constitutes an important sector of the industry. Operators exploiting such deposits are concerned with the most appropriate mining method and equipment to mine in such a way as to minimize waste rock dilution. Unfortunately, most of the equipment available does not meet the additional geological and economical restrictions typical of narrow vein mining. Thus, mining these deposits is labor intensive and can result in low productivity. Consequently, the profitability of many such mines is reduced. Increased mechanization is one of the solutions to ensure their healthy survival. This paper reviews the state-of-the art in mining equipment with reference to narrow vein mining limitations. Problems encountered Geological conditions and economical factors have a decisive impact on the choice of mining method and equipment. Many ore bodies are relatively small, usually erratically distributed, often less than 2 m (6.5 ft) wide, with dips ranging from 0° to 90°. To make things worse, ore body limits are sometimes difficult to determine and the ground is hard and abrasive.

Jan 1, 1992

By G. Biles, K. Loughran, R. Anand

The Eskay Creek Mine, a high-grade gold and silver underground mine in northwestern British Columbia started operations in late 1994. For the first three years of operation, ore was crushed to minus 2 inch; blended to satisfy smelter terms and direct shipped to smelters in Japan and Canada. Before commencing the above operations, preliminary evaluations had indicated that an extremely complex metallurgical facility, requiring a large capital investment, would be required to recover the precious and other economic metals. Subsequently markets for raw ore were identified and underground mining and ore transportation operations, at a nominal production rate of 300 tonnes per day, were commenced. Exploration diamond drilling in 1995 and 1996 discovered additional lower grade mineralised zones, which are less economic or uneconomic as direct shipping ore. Test work demonstrated that these lower grade zones could be upgraded to marketable concentrates in a simple flotation mill. The mill was built in 1997 and has been in operation since late 1997. The mining rate now has increased to 400 tonnes per day including 250 tonnes per day of the direct shipping ore and 150tonnes per day of the mill ore. Prime Resources Group Inc. wholly owns the Eskay Creek Mine. Homestake Canada Inc. owns 50.6% of Prime’s shares and manages and operates the mine on Prime’s behalf. This paper discusses the mining and milling operations as well as ore shipment and smelter contracts' administration.

Jan 1, 1999