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By Jerome C. Motz

Los tipos, tamaños y número de trituradores empleados en un sistema completo de reducción varían con factores tales como: el volumen de mineral que debe ser procesado, tamaño de la mina, tamaño máximo del todo-uno, dureza del mineral y de su roca soporte y tamano requerido para el producto final. Básicamente, sólo existen tres tipos distintos de equipos de trituración para la etapa primaria. Estos son: machacadoras de mandíbulas, trituradores giratorios y molinos de impactos, y cada uno tiene unas características operativas propias. Las trituradoras de mandíbulas operan por aplastamiento de las rocas entre dos paredes, una fija y otra móvil, en una cavidad determinada. Han sido ensayadas variaciones en la caída y movimiento, pero la mayoría de las máquinas del tipo Blake tienen un ángulo de trituración de aproximadamente 27° entre las mandíbulas fijas y móviles. El principio de trabajo del triturador giratorio es muy simple. Si un cono es montado en un eje vertical y la parte superior del eje se sitúa estacionaria mientras que la parte inferior gira excéntricamente, el cono tendrá también un movimiento excéntrico. Si el cono se coloca en una carcasa conveniente se movers aproximándose y retrocediendo de las paredes que lo contienen a la vez que gira. Si el cono y las paredes de la carcasa son suficientemente robustas y pesadas, cualquier cosa apresada entre ellas será triturada. Al lado de los incrementos de la eficacia desarrollados por las continuas acciones de trituración y el curvado de las caras de trituración, los trituradores del tipo giratorio tienen las siguientes ventajas: 1. La gran abertura de alimentación disponibles en comparación con trituradores de otro tipo. 2. El alto rango de tamaños y capacidades, que se encuentran entre 600 y 6.000 toneladas por hora. 3. Generalmente no es necesario control en la alimentación. Dumpers de 150 a 200 toneladas descargan directamente en las actuales instalaciones.

Jan 1, 1982

By Charles A. Stone, Eric C. Wang

"Top-down wall construction was selected to excavate the tunnel portal headwall areas of the ORB East End Crossing tunnel. Benefits included an expedited construction schedule realized by staging excavation and rock support with development of temporary construction access roads. Rock mass properties, proximity to right-of-way lines, and blasting impacts were considered to design suitable lift heights, intermediate bench locations, and permanent ground support systems to provide supported permanent excavations. Carbonate rock formations contained adverse subsurface geotechnical features including solution cavities, clay seams, weathered joints, pinnacled rock, and boulders, which were addressed by flexibility in design. The resulting over-break conditions and long-term water infiltration issues required economic design solutions. This paper discusses the unique considerations required during design of top-down tunnel portal headwalls excavated in carbonate rock formations including temporary excavation and permanent ground support systems which provided flexibility for unforeseen over-break, minimized construction delays, and provided long-term stability. PROJECT BACKGROUND The East End Crossing is an approximate 8-mile highway connection of the Louisville-Southern Indiana Ohio River Bridges East End Crossing (ORBEE) project. Currently under construction, this connection will link State Road 265 in Indiana with the Gene Snyder Freeway in Kentucky to form an outer loop bypassing the Louisville-Southern Indiana metropolitan area. This highway project features a 2,510-foot long, three-span cable-stayed bridge over the Ohio River, and a 1,680-foot long tunnel. The project, being managed by the Indiana Finance Authority and Indiana Department of Transportation, is scheduled to open to traffic by the end of 2016. The new twin three-lane tunnel under the historic Drumanard Estate requires steep cut excavations for approach and tunnel portal walls on the order of 60 feet deep through overburden soil and carbonate bedrock formations. This paper focuses on the design considerations for the permanent support of the top-down wall construction of the tunnel portal headwalls and approach retaining walls."

Jan 1, 2016

By S. Dereski, G. Chancellor

"INTRODUCTION Barrick Goldstrike underground mine is located in Northern Nevada on the Carlin trend. The underground mine produces approximately 1.5 to 1.8 million tons a year and approximately 400,000 ounces of gold a year. The mine is comprised of two main mining methods, underhand drift and fill and long hole stopeing. The mass blast project stated in 2012 as Value in Use project between Barrick, Southwest Energy and Orica, the goal of this project is to be able to blast an entire drilled stope in one blasting event instead of taking multiple blasts. This will improve productivity by eliminating intermediate blasts, reducing additional blast time, eliminate drill hole cleaning due to the previous blast. Project implantation started in the fourth quarter of 2013 and mass blasting has been common practice throughout all of 2014. STOPPING PRACTICE AT GOLDSTRIKE BEFORE MASS BLAST PROJECT The production from the Goldstrike mine is approximately 70% from stopes. The timely completion of all stoping blocks is imperative to maintaining the plan, delays in the execution of any particular block can have detrimental ripple down effects as neighbouring stoping blocks become delayed. Before the implementation of this project the standard practice for Goldstrike was to: 1. Drill the entire stope 2. Blast a bottom shot, normally between 10 feet and 20 feet off the bottom of the slot rings. This would be mucked out before the next shot 3. Blast a pull through shot, this included the raise and one or two production rings. Again mucking this out before the next shot 4. Blast production rings, one or two more blasts depending on stope size. In most cases between each of these blasts the holes in the other rings of the stope would have to be cleaned. This meant either cleaning manually using air and water, or mechanically cleaning with a bench drill. In many cases this resulted in delays in completing the stoping block."

Jan 1, 2015

By Bruce K. Cantrell, Stephen W. Stein, Kenneth L. Williams, David Hassel, Harvey Patashnick

In June 1992, the Mine Safety and Health Administration (MSHA) published the 'Report of the Coal Mine Respirable Dust Task Group; Review of the Program to Control Respirable Coal Mine Dust in the United States." As one of its recommendations, the report called for the accelerated development of two mine dust monitors: (1) a fixed-site monitor capable of providing continuous information on dust levels to the miner, mine operator, and to MSHA, if necessary, and (2) a personal sampling device capable of providing both a short-term personal exposure measurement as well as a full-shift measurement. In response to this recommendation, the National Institute for Occupational Safety and Health (NIOSH), Pittsburgh Research Center, initiated the development of a fixed-site machine-mounted continuous respirable dust monitor. The technology chosen for monitor development is the Rupprecht and Patashnick Co., Inc. tapered element oscillating microbalance. Laboratory and in- mine tests have indicated that, with modification, this sensor can meet the humidity and vibration requirements for underground coal mine use. NIOSH is continuing that effort by developing prototypes of a continuous dust monitor based on this technology. These prototypes are being evaluated in underground coal mines as they become available. This effort, conducted as a joint venture with MSHA, is nearing completion with every promise of success. The immediate benefit of this effort will be to researchers, regulatory personnel, and mine personnel, by permitting evaluation of specific mining practices to see which expose mine workers to excessive dust levels. Using this information, mine personnel can optimize mining procedures to reduce dust exposure. MSHA and the operators will also be able to use the dust concentration data to judge whether dust plan parameters are adequate to continuously maintaining environmental dust levels below the applicable standard. The second development recommended by the Dust Task Group was a person-wearable version of the continuous dust monitor. It will be a monitoring system designed to provide a measurement of worker exposure to respirable dust, both during and at the end of the shift.

Jan 1, 1997

By B. Ollila, L. Brown

As mining fronts converge in sill pillar extraction, the mine seismic hazard can be significantly altered and large seismic events have been known to occur. This study used common seismological tools to analyze the extraction of highly stressed sills from two deep Canadian mines with the objective of characterizing the rock mass failure process and relation to large seismic events. The approach treated a sill pillar as a mine-scale failure process, meaning the temporal and spatial analysis scales needed to exist on a similar scale to a significant part of the mine. The temporal analysis scale overlapped with the entire mining sequence, included several years of data to establish long term change in the characteristics of rock mass failure. The spatial analysis scale included the rock mass surrounding a large part of the convergent mining sequence. To accommodate the large spatial and temporal scales, the highest magnitude 2% of events in the dataset were analyzed. Similar results were shown for sill pillars in both mines. Changes in event location, time, magnitude, energy index and apparent volume were similar to expected behavior of rock failure. Indications of strain hardening and strain softening were observed in the sills. The hardening period was characterized by a high rate of small events, high energy index and low apparent volume, giving an indication of high stress and low rock yield. The softening phase was characterized by a decrease in the rate of small events concurrent with an increase in the rate of large events, low energy index and high apparent volume, indicative of release of stress and increase in deformation. Large magnitude seismicity was strongly related to the softening phase, with several Nuttli magnitude 3.0+ events occurring shortly after the transition from hardening to softening failure regimes. Sill failure was indicated by consistently low EI and low levels of seismicity and seismic hazard in the sill pillar.

Apr 26, 2022

By J. Fuller, J. Beale, Q. Justice, Z. Agioutantis, S. Hicks, Z. Khademian

Seismic events can be induced by mining due to unstable rock failures or slips along geologic structures leading to the sudden release of strain energy stored within the rock near underground openings. Depending on their intensity and proximity to the working area, seismic events may raise safety concerns, and thus the forecasting of the seismic potential in advance of mining can mitigate safety hazards. The integration of seismic data into geomechanical modeling can provide a tool for identifying areas with elevated potential for seismic activity. In previous works, a methodology was developed to forecast seismic potentials in a pseudo-2D geomechanical model. In this paper, seismic events, ranging in magnitude from about -1.0 ML to 3.7 ML, recorded in a deep longwall mine (> 450 m) in Virginia during mining of six adjacent panels, were evaluated to identify areas where larger magnitude events (> 1 ML) had occurred. A three-dimensional geomechanical model of the same area was constructed in the 3DEC software to calculate the state of energy in the system with the progress of mining along the entire length of each panel. The variation in the energy content of the model was monitored, and zones with potentially damaging slip-type events along bedding planes followed by compressive-type events in the sandstone roof are the mechanism for large seismic events in the studied mine. The locations of anomalous events corresponded to the regions where both compression-type SPI (C-SPI) and slip-type SPI (S-SPI) were elevated. The modeling results were compared with the recorded seismic events to evaluate the model performance in forecasting the seismic potentials. The model was able to forecast 100% of regions with elevated and moderate seismic potential and 89% of regions with low seismic potential. This shows the applicability of energy balance calculations for forecasting the approximate location and timing of destructive seismic events in advance of mining.

Jun 25, 2024

By David A. Lindsey

The world-famous beryllium deposits at Spor Mountain, northwest of Delta, Utah, are the principal resource of beryllium in the United States. Beryllium metal, beryllium-metal alloys, and beryllium oxide ceramics have a long history of specialty applications. Beryllium is used in X-ray windows, nuclear reactors: and a wide variety of industrial and electronic products. The Spor Mountain deposits are located on the western rim of the Thomas caldera, one of at least three volcanic depressions that subsided during voluminous eruption of rhyolitic ash-flow tuff in early Oligocene time. Beginning in early Miocene time, the caldera rim and its volcanic fill were cut extensively by basin-and-range faults. Basin-and-range faulting was accompanied by local eruption of high-silica rhyolite. High-silica topaz rhyolites at Spor Mountain (21 Ma) and Topaz Mountain (6 to 7 Ma) are especially rich in fluorine and lithophile trace elements such as beryllium. Topaz rhyolite of the Miocene Spor Mountain Formation, or its magma was the source of the Spor Mountain beryllium deposits. At Spor Mountain, deposits of beryllium, as submicroscopic bertrandite, Be4Si207(OH)2, occur in rhyolite tuff beneath flows and domes of topaz rhyolite. The beryllium deposits also contain minor disseminated fluorite, lithium, uranium, and anomalous traces of other lithophile elements. The host tuff, originally composed of glass, zeolite, and abundant clasts of carbonate rock, has been altered to clay and potassium feldspar. Fluorite, quartz, and opal replaced carbonate clasts in the tuff. Beryllium was deposited with fluorite in altered clasts and in the matrix of altered tuff. Fluorspar and uranium deposits are spatially and genetically associated with the beryllium deposits at Spor Mountain. Fluorspar deposits at Spor Mountain occur mostly in breccia pipes. Some fluorspar pipes contain uranium, thought to be dispersed in the structure of fluorite. Secondary uranium minerals are dispersed throughout the tuff that hosts beryllium deposits and comprise lenses of ore-grade uranium at the Yellow Chief mine.

Jan 1, 2001

By A. Wahab Khair

There are a number of factors, including mine geometry, local geology, geotechnical properties of overburden materials, surface topography, sequence of overburden strata, and extraction rate, affecting characteristics of surface subsidence imposed by large underground excavation. Among the factors, the first three have the major impact on magnitude and area of influence of final subsidence. Hence, a model has been developed, considering mine geometry and geologic and geotechnical properties of the associated area, to predict maximum possible subsidence and subsidence profile at any cross-section within the subsidence influence zone. Based on extensive field measurements and mechanics of strata behavior, under the action of overburden pressure in the vicinity of the large underground excavation, the vertical plane of influence zone has been divided into five areas: 1) excavated, 2) caved, 3) fractured, 4) highly deformed but continuous, and 5) subsided zones. In the final state of subsidence each area maintains a certain volume expansion/contraction, i.e. a factor which depends on geologic, geotechnical properties of overburden strata and state of stress within the subsidence trough. The cumulative volume expansion and contraction partially compensates for the extracted volume and the difference between the two is reflected as the subsided area in the form of the area within a curve defined by a probability function. The input parameters for the model are mine geometry (i.e. length and width of the panel), depth of cover, and geologic and geotechnical properties of overburden strata (i.e. percentage of hardrock, sandstone and limestone, and soft rock, shale within the overburden strata and Rock Quality Designation). The model calculates angle of draw for the site, expansion factor for each zone utilizing empirical relationships, maximum possible subsidence and plots the subsidence profile curve. The model is developed for a longwall excavation with unique characteristics of prediction of angle of draw and maximum possible subsidence for a given geologic and mining condition at any cross-section within the influence zone of subsidence.

Jan 1, 1986

By David A. Summers, Clark R. Barker

INTRODUCTION The use of high pressure water as a means of excavating valuable mineral from the earth dates back many hundreds of years, but is most familiar through its use in sluicing the gold bearing deposits of California and Alaska. Interestingly, that practice, which dates back from at least 1852, is now nearly defunct with only one known surviving hydraulicking operation.(l) In that particular development water under a relatively low, natural pressure head is carried to a nozzle and from the nozzle it is thrown a distance of perhaps 20 ft to wash the gold bearing material down into a channel from whence it's carried over some mineral collector, and the water is then recirculated. Development of this practice in coal mining began in Russia in 1915, and Dr. Muchnik established the criteria for hydraulic mining during the 1930's. The first water jet mine was established underground in 1939 and has been followed by several larger mines to the present day where there are at least 11 large mines producing about 10 million tons of coal per year in the Soviet Union.(2) Water jet mining has expanded from Russia through Japan, Germany, Poland, Czekoslovakia, and mainland China. The nearest location to the United States is currently at the Kaiser Mine in Sparwood, British Columbia, where a three-man crew has mined up to 3,400 tons of coal in a 6-hr shift. There are many advantages to this particular operation. It is one of the safest mining operations for coal in Canada and has won safety awards because of this. The process is relatively cheap since the cost of the equipment is much lower than that which would be required conventionally using either longwall or shortwall mining, and transportation is achieved by fluming the coal to a central station using a gravity feed and then pumping to the surface, (3) For this

Jan 1, 1976

By Henryk Bystron

The underground fire areas are formed as a result of failure of an active fire extinguishing in mines. Such areas bring about a fire hazard for the work safety and for the mining operations. For liquidation of fire areas it is necessary to create conditions under which a possibly small amount of air is flowing to the fire. To achieve this end, isolation methods are used which consist in erection of full stoppings of various types, sealing practically the fire areas. These methods usually fail in oases when the coal seam is thick and the fire area is situated between the intake shaft and the outtake shaft in which a high-pressure fan is in operation. In such oases ventilation methods based on the balance of aerodynamic potentials in fire areas are to be used. For a very long time the isolation and ventilation methods have been developed in the world mining, No sooner than since 1980 the aerodynamic potential has been used in ventilation methods [1,2,6]. So far, the aerodynamic potential was called in different ways and was determined by different formulae which gave its numerical values more or less approximative. Nevertheless, even in situations featuring by a lack of sufficient theoretical bases of the aerodynamic potential the ventilation methods brought considerable profits in the field of safety in mines and of economics in mine ventilation [1-6,9,10]. The increase in the depth of underground mines and their equipment with expensive machines (apparatus) brought about that the Central Mining Institute at Katowice undertook research works which are crowned by the theory of the aerodynamic potential in mine ventilation networks [7]. The use of this theory in the mentioned ventilation methods will increase their effectiveness. Since the theory will be published in the III rd quarter of this year, we cite in the present paper the formulae and relations associated with this theory which are indispensable to present the balance of aerodynamic potentials in the fire areas in their control by ventilation methods.

Jan 1, 1980

By L. Xu, R. Bhaskar

The trends in dust exposure levels and longwall production over the last 20 years were examined. The study was conducted for nine major underground coal mining states: Alabama, Colorado, Illinois, Kentucky, Ohio. Pennsylvania, Utah, Virginia and West Virginia. The dust exposure and shift production data were obtained from the Mine Safety and Health Administration (MSHA). Trends were identified and analyzed for the headgate, tailgate, jack setter and shearer operator positions. The relationships between average scam height and dust exposure levels as well as the percentage of dust samples exceeding the dust standard for the four occupations were examined. The concentration to production ratio, a better indicator of the success of dust control measures, was calculated for the 20-Year period. The data show that, in general, dust exposures over the last two decades hare decreased. However, it appears that much of the decrease was in states that had significantly high dust exposure levels (such as Alabama and Utah), while for states such as Virginia, West Virginia and Pennsylvania, the decrease was not as dramatic. Major reductions were made in the period starting in 1982, when a host of longwall dust control techniques was introduced. However, nearly 35% of jack setter and shearer operator dust exposures were still above 2 mg/m3 in 1990, with nearly 20% of tailgate operator exposures and 8% of headgate operator exposures being above the federal standard. The concentration/production ratio has shown a remarkable decrease since 1981. This indicates that long wall operators have been able to meet increased production as well as, to a large extent, control chest production at the face. An analysis of the relationship between average seam height and dust concentration for the years 1984 and 1990 shows that while in 1984, the dust concentration could be correlated to seam height, in 1990 the relationship was less definitive. This indicates that technological factors may be more important than seam height.

Jan 1, 1994

By Jeong-Heum Kim, Jan M. Mutmansky

The primary purpose of the research summarized in this paper was to experimentally investigate the performance of numerous mine ventilation network iterative solution procedures under controlled conditions to determine which procedures were most favorable in terms of number of iterations and total computational speed. The solution methods investigated included: (1) the Newton-Raphson method, (2) the Hardy Cross method, (3) the linear theory method, and (4) the second-order approximation method. The first three of these methods have been widely used while the fourth was formulated by the authors to investigate its potential for ventilation network computation. In addition to the four solution procedures, three methods for selecting the branches in the meshes and three methods for ordering the meshes for computational purposes were also studied during the investigation. As a result of these various options, a total of 36 combinations of solution methods, mesh selection procedures, and mesh ordering schemes were studied. To investigate the performance characteristics of these combinations, an identical set of 180 randomly chosen ventilation networks was used in sets of 30 networks at six different levels of complexity. Here, the level of complexity is defined as a function of the size (number of branches) and the density (number of nodes compared with the number of branches) of the network. To eliminate variations in the solution times due to differences arising from the fans, a single fan with a fixed characteristic was utilized in all the experiments. This paper will deal only with the results for the four iterative methods used to achieve solutions to the mine ventilation networks. The minimum-resistance spanning tree method for mesh selection and random mesh ordering was used throughout the analysis. The conclusions will be based upon a detailed accounting of the efficiency of each method using the total time to achieve a solution to each problem and averaged over the total problem set. However, the number of iterations and the time to complete each iteration will be studied in order to better understand the solution procedures.

Jan 1, 1991

By Steven L. Webb

In 1965, Monsanto started planning of a new dike (No. 15) for use in disposing of ore tailings at its Columbia, Tennessee phosphorus plant. Generally, in the phosphorus industry only clay tailings are produced making it necessary to borrow fill for construction of a retaining dike. Borrow area investigations indicated that weathered shales and limestones adjacent to the dam site would provide adequate construction materials. It was proposed that the more clayey weathered shales would be used for the impervious core and the more granular weathered limestones would be placed in the outer shell of the dike. A dike section is shown in Figure 1 with gradation curves for core and shell materials in Figure 2. Figure 3 shows an aerial view of the tailings pond and dike during first-stage construction. Although the borrow area explorations indicated that the weathering of the sedimen-tary deposits was highly irregular, it was not considered necessary, at the time, to specify special excavation techniques that would ensure a thorough blending of the materials in the borrow pit. Instead, because of the short haul distance, a scraper operation for both clay core and granular shell materials was adopted as the most economical construction method. Using this materials-handling technique, variations in permeability could be expected in the compacted layers inthe embankment. Where continuous layers of less clayey limestones were placed in the core of the dam, extending from the upstream to downstream shells, seepage losses would be expected to increase substantially. Upon completion of the damn 1966, and shortly after reservoir filling, it was apparent that the construction technique used had resulted in the formation of pervious layers through the clay core. Excessive seepage on the downstream face of the dike was noticeable within a very short time. A monitoring program was established, using piezometric observations and flow measurements. These data indicated that seepage should be reduced and the application of an asphaltic emulsion

Jan 1, 1971

By Jonathan M. Pachter

Introduction The Bailey mining complex is located in Greene County, PA, in the Talley Run valley. It includes two Pittsburgh No. 8 seam deep mines (Bailey and Bailey No. 2), a coal preparation plant, and two refuse areas. Clean coal is shipped from the complex by way of a 24-km (15-mile) rail spur. It con¬nects to the Monongahela Railroad in Waynesburg, PA. In early 1981, Consolidation Coal Co. (Consol) began a permitting effort aimed at building the mine. Construction began in Dec. 1981, less than six months after the first application submittal. Clean coal was produced by the Bailey preparation plant in Sept. 1984. Several factors allowed quick permitting at Bailey. Effective project planning, including early formulation of a project planning team, and identification of red flag issues helped. So did innovative responses to environmental questions, and development of environmentally sound project designs. Forming a project planning team Consol management decided that Bailey mine construction had to begin in late 1981. A project permitting team was immediately formed to see that this goal was achieved. Team members included a project manager and personnel from Consol's engineering, environmental, and operating departments. Communications among all team members was crucial in obtaining the various state and Federal environmental permits. All project permitting team members participated in weekly meetings during application preparation. Discussions at these meetings ranged from recent public contacts, to application content, to environmental design considerations. Establishing a project manager was critical to the team's success. The project manager was responsible for seeing that permit acquisition, facility construction, and operation of the mining complex proceeded on schedule. The project manager had the final word, mediated disagreements, and frequently served as project spokesman to the public. Identifying red flag issues Red flag issues are those environmental factors with the most potential for delaying the schedule or stopping the project. The first red flag issue identified was organized local opposition. This opposition had completely derailed a US Soil Conservation Service proposal to build a multipurpose dam on Enlow Fork, a short distance

Jan 9, 1985

By G. L. Bloomsburg, E. G. Zahl

INTRODUCTION During the past decade, disposal of mine and mill waste material has become an increasingly important issue. Problems of structural stability and seepage effects on stability have been well addressed, and significant technological advances have been achieved [(1-4)]. The primary problem currently facing industry in the waste disposal area is environmental protection. The greatest environmental concern is over the amount of seepage or leachate entering the ground or surface water systems. Methods to determine the quantity of water seeping from a tailings pond have also been covered in the literature [(5-23)]. However, much further work is necessary to fully implement these techniques to evaluate tailings pond performance and accurately predict the various influences of disposal management on seepage from mill tailings. This study is an effort to develop those necessary techniques. The partially saturated method of analysis which is discussed here is applicable to most situations involving seepage from waste piles. In a conventional type of tailings pond, a slime layer is ordinarily built up under the ponded area. This material will consolidate slowly from capillary pressures and the weight of material deposited above it. In situations where this consolidation results in a slime layer of lower permeability than the native foundation soil, a partially saturated flow condition will usually exist in the native soil. This will result in a smaller quantity of water entering the ground-water system than would be predicted by saturated flow analysis. The finite-element program used in this study was originally developed by Neuman [(17)] and is called UNSAT2. The program will analyze problems of saturated or unsaturated, steady or nonsteady flow and can be used with a variety of boundary conditions such as constant flux, constant pressure, maximum flux, minimum pressure, and with infiltration, evaporation, and seepage surfaces. The maximum conditions are controlled by exterior factors such as atmospheric conditions; but the actual flow in the porous media is determined by properties of the porous media. In this test, the computer code was used

Jan 1, 1980

By Jerzy F. Sobala

Large - scale tests at the Polish Safety in Mines Institute /established in 1925 as "Barbara" Experimental mine/ performed in underground and surface experimental galleries, simulating conditions in roadways in mines, are the basic source of knowledge about development and propagation of explosions. Investigations up to 1973 were performed under the Institute's former research director Professor W.B. Cybulski s supervision, whose handbook concerning combating of coal dust explosions in collieries has been also published in the U.S. (2), and are continued until now. For investigating explosions the "Barbara" Mine disposes of 170 m /7.5 sq m/, 200 m /5 sq m/ and 400 m /7.5 sq m/ underground gallieries as well as of 10 m, 25 m, 44 m and 100 m surface galleries. For example, in the 200 m gallery more than 5200, and in the 400 m gallery - more than 2200 test explosions have been performed. Basing on determination of explosion hazard in roadways in collieries and on data of scientific research performed at the Safety in Mines Institute, the strategy of combating the explosion hazard in mines was elaborated; it includes mainly diminishing of coal duet. deposition in roadways, and -because of the possible inefficiency of this safety means in some roadways in mechanized collieries -- the introduction of three additional defence lines against coal dust explosions. It should be mentioned that at the "Barbara" Mine the problem of coal dust explosion initiation was elaborated as contract work for the U.S. Bureau of Mines. The main aim of this paper is to discuss some aspects of methods and measures for combating coal dust explosion hazard in highly mechanized mines in the Polish mining industry where in some roadways explosible coal dust is present, but it is not the aim to evaluate the effectiveness of methods and measures for diminishing the make-up of coal dust and coal dust deposition rates in roadways.

Jan 1, 1980

By Ralf Winterberg, Axel G. Nitschke

"Macro synthetic fiber reinforced concrete or shotcrete is seen by many design engineers as offering a viable alternative to steel reinforcement in tunnel linings. The technology is now commonplace for primary and permanent ground support in both mining and civil tunnel applications. It has for instance become the standard form of reinforcement in the Australian mining industry, and has been used for a majority of permanent tunnel linings in recent tunnel construction in Norway. Similarly, macro synthetic fibers are becoming a standard solution for the initial lining in the USA.The use of macro synthetic fiber offers innovative solutions, yielding robust and sustainable tunnel lining designs. Citing recent research and actual projects, this paper presents state-of-the-art design considerations for fiber reinforced tunnel linings relating to structural and long-term performance. Topics include seismic resistance, crack width control, corrosion and durability, as well as sustainment of performance with age.INTRODUCTIONMacro synthetic fiber reinforced concrete (MSFRC) and shotcrete (MSFRS) has reached maturity as an engineered material and is widely used in all forms of tunnel linings, for temporary as well as permanent ground support in both, mining and civil tunnel applications. In the Australian underground mining industry, it became the standard form of reinforcement, where 2014 literally marked the end of steel fiber use in shotcrete [1], and has been used for over 80 percent of permanent tunnel linings in recent tunnel construction in Norway. Similarly, macro synthetic fibers are becoming a standard solution for tunnel initial linings in the USA. Recent examples in civil tunneling are the Devil’s Slide Tunnel and Caldecott Fourth Bore Tunnel in California or the Anacostia River Tunnel Intershaft Connector Tunnel in Washington D.C. In addition, an increasing number of tunnels are adopting shotcrete permanent linings using macro synthetic fibers. Examples include the A3 Hindhead tunnel near Guildford in the UK and the North Strathfield Rail Underpass in Sydney, Australia [2, 3]."

Jan 1, 2016

By Hans Ehdwall, Erling Strandon, Friedrich Steinhäusler, Victor A. Leach

INTRODUCTION The basis for any measures in radiation protection is the quantitative assessment of the radiation exposure. Due to the extensive requirement in health physics manpower and related cost it is necessary to develop certain criteria regarding the necessity of monitoring and the selection of population groups to be monitored. Depending on the overall objective of the particular programme, the specific requirements for the measurement of atmospheric levels of radon, thoron and their daughters for practical radiation protection purposes differ significantly for those in the case of research applications, such as use of data for epidemiological studies. As part of the OECD-NEA Programme on Radon Dosimetry and Monitoring Task Group 1 reviewed available models for lung dosimetry in the case of inhalation of radon 222, radon 220 and their daughters (James et al., 1981). Using defined reference atmospheres for underground- and open-pit mining conditions as well as for dwellings, normalized conversion factors were calculated for the mean bronchial dose (mGy) per Working Level Month exposure (WLM). Task Group 2 discussed the features and adequacy of the Working Level (WL)-concept as a physical correlative of the absorbed lung dose and its limits of confidence and accuracy (Johnson and Leach, 1981). The WL-concept was found acceptable for radiation protection purposes provided that modifying factors like particle size, unattached fraction and respiratory minute volume did not cause the effective dose equivalent per WLM as calculated by either the James-Birchall or Jacobi-Eisfeld lung models to be more than 50 % above that given by ICRP of 0.01 Sv/WLM. For application in epidemiological studies the WLconcept is considered acceptable if the calculated absorbed dose to the tissue does not vary by more than 50 % (at the 95 % confidence level) from that given for defined reference conditions. In the following, objectives and requirements for the measurement of radon, thoron and their daughters are discussed for operational health physics routine and special research applications according to the scheme outlined in Fig. 1.

Jan 1, 1981

By Yi Luo, FENGLIANG WU

Explosions and fires originated from longwall gob due to the formation of methane-air mixture have been a severe threat to coal miner’s lives.Many numerical studies on coalmine fire and explosion hazards have focused on the airflow in roadways and mine gobs. However, most of these studies isolate the gob from its surrounding roadways, and the network analysis and the CFD method are applied independently tomodel the two classes of airflows. This approach greatly limits the ability of simulatingmine ventilation flow, especially unable to consider the effects of gob boundary conditions on air exchange between the gob and the surrounding airways. An innovative finite tube method (FTM) is developed to couple the one-dimensional mine ventilation network (MVN) and the 2D/3D gob flow field (GFF). In FTM, GFF is discretized into a finite number of flow tubes each of which is formed by any two adjacent stream lines. These tubes, representing the gob’s field flow, connect the MVN into a new coupling network. To solve the coupling model betweenMVN and GFF, an iterative solution technique is developed inwhich the MVN analysis is used to evaluate the boundary pressures for GFF simulation and in turn the FTM feeds the GFF results back to the coupling network. Based on the FTM approach, a model for gas migration in gob has been established for delineating the hazard zones of explosive methane concentration and spontaneous combustion. A computer program is developed to implement the FTM simulation. An illustrative example with five flow tubes representing the GFF is created to verify the stability and convergence of the FTM solution process. A simulation example also indicates that the accuracy of FTM is improved by 12% compared with previous method. Results of an application case show that the program is capable of quantitatively evaluating the gob’s risk zones prone for spontaneous combustion and gas explosion as well as performing risk analysis for various ventilation scenarios.

Jul 10, 2020

By Huayang Luo, Hui Li, Wei Zhou, Gang Chen, Scott Huang

Abundant beach and offshore resources of gold deposits are demonstrated along the coast of Alaska, and Nome is one of the richest places. Considerable amounts of mineral resources data were collected and stored in geodatabase of ArcGIS. In searching for the best model for the interpretation of the Nome offshore gold placer deposit and accurate estimation of the reserve, the data were processed and visualized using ArcGIS. The gold concentration data collected from the Nome offshore deposit are of characteristics such that: (1) near lognormal distribution of data, (2) high positive skewed distribution, (3) a small number of extreme values dominating the reserve estimation, and (4) apparent anisotropy of the variogram. These characteristics make the development of an interpolation model for the gold deposit a challenging task. In this study, it is determined that five geostatistical methods provided by ArcGIS with visualization capability, i.e., Inverse Distance Weighted (IDW), Ordinary Kriging (OK), Ordinary Kriging with lognormal transformation (OK-log), Simple Kriging (SK) and Indicator Kriging (IK) are suitable candidates for the development of an interpolation model. The five algorithms are evaluated and compared with each other in the study. Cross-validation technique is employed to examine the accuracy of the five approaches. It is determined that the OK and SK algorithms are not suitable for this dataset because of the difficulty of fitting the semivariogram model despite their lower RMSE (Root-mean-square error) and the higher RMSS (Root-mean-square Standardized). The OK-log algorithm does not provide satisfactory prediction either, because it places too much weight on the role of extreme values resulting in overestimation and extremely high RMSE. Reserve estimation of each blocks is conducted using each of the interpolation methods and compared with the reserve calculated using the conventional polygon method. The IWD and IK algorithm seem to provide estimations more agreeable with the conventional polygon method.

Jan 1, 2005