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By J. Cheung, H. Rabia

The traditional methods for predicting mine air temperatures are based on steady state heat transfer models as developed by Voss (1969) and Starfield (1983). These models are useful in providing approximate predictions of air temperatures in trunk and gate roadways. More complex models are required to provide predictions along faces. In practice, the heat flow in underground mines is of transient nature and recent observations in some British Coal mines show temperature variations of up to 5°C during a working week at any one point. In addition, the recent introduction of district recirculation of mine air in British mines requires a more precise prediction of air temperatures during start up of the system or during breakdown. A transient heat transfer theory based on the work of Jordon is presented. The method lends itself easily to computers due to the large number of calculations involved. The proposed transient model is validated using actual temperature data collected in roadways in Wearmouth Colliery in the North East of England. The transient results are compared with the steady state predictions of Voss (1969). Finally, predictions of air temperature in roadways for a recirculation system will be made using both the transient and steady state models. Temperature predictions along faces will be made using a separate model.

Jan 1, 1989

By C. Newman, N. Schaefer

"As underground mining continues to operate at deeper depths and in more complex geologic and geometric conditions, there has been an increased demand for mine designs to better reflect site-specific conditions, measurements, and observations with respect to ground control safety and stability. During the past 20 years, the National Institute for Occupational Safety and Health (NIOSH) has developed, validated and distributed a number of free software packages (ALPS, ARMPS, AMSS, ARBS, etc.) specifically aimed at providing the industry with ground control design tools for the evaluation of underground stability in coal mines. While these programs were initially developed as stand-alone software packages where all data reside on an individual or workgroup computer, modern day web applications have begun to migrate toward a “multiple-tier” architecture, allowing multiple users to access an application by means of an Internet connection.This article introduces a new Web-based platform for ground control software applications, webGroundControl (webGC). webGC provides users with a platform through which one can access NIOSH and other ground control design software utilizing a Webbased, multiple-tier architecture. This provides users with easy access to ground control designs, the ability to perform on-demand calculations, as well as increase online collaboration between operational personnel and engineers through a web server. The basic framework and, more specifically, the user-interface of webGC are defined and outlined within this paper. It is expected that the mining industry will rapidly embrace this new product and webGroundControl will be a new paradigm for mining engineering computer applications. Platform architectureAs Web-based computing technologies continue to advance and cloud computing gains momentum, web-based platforms for software applications are growing in popularity across all industries due to ease of product accessibility and distribution as well as data security. With the expansion of Internet connectivity on the surface and underground, there is an industry need to develop Internet-based applications for mine and ground control design (Barczak et al., 2016). Cloud-based products are already available for many industries and are slowly being developed for the mining and geotechnical industry. A notable example is the drillysis package, which is a cloud-based borehole logging database (Geosysta, 2017) or the Dynonobel package for on the fly blasting calculations (Dyno Nobel, 2013)."

Jan 10, 2017

By David M. Sullivan

Jim Walter Resources, Inc., (JWR) a division of Jim Walter Corporation, has 5 underground coal mines in Central Alabama varying in depths from 400-2500 feet. Gas Chromatography plays a vital role in safety and production at each mine. Chromatographic data is used to monitor working areas for safety and proper ventilation and to control the quality of extracted methane for sales purposes. Instruments have a1 so been employed during mine rescue operations. JWR's first Gas Chromatograph (GC) was installed in April 1984 to monitor gob gas taken from longwall gob areas and sold based upon BTU value. The GC allows maximum extraction rate of the gas within the quality guidelines. By late 1985, another system was purchased and both units were upgraded to perform an analysis of indicator gases which may appear during mine emergencies. Each unit has the ability to detect and measure the following gases at almost any level ; Oxygen (021, Nitrogen (N2), Methane (CH4), Carbon Dioxide (C02), Hydrogen (HZ), Carbon Monoxide (CO) , Ethylene (C2H4), Ethane (C2H61, Acetylene (C2H2) and a composite of hydrocarbon threes and above (C3+). The analytical run is accomplished in less than ten minutes and is completely automated. This report will introduce Gas Chromatography as applied by Jim Walter Resources for the analysis of underground mine air.

Jan 1, 1987

By C. R. Stephan, N. B. Greninger, W. A. Slivensky

The U.S. Bureau of Mines and the U.S. Mine Safety and Health Administration (MSHA) are participating jointly in a research program to evaluate strength characteristics and air-leakage resistance of various proposed seal designs for use in underground coal mines. Over 20 full-scale seals were constructed in the Bureau's experimental mine at the Lake Lynn Laboratory, air-leakage tested, then subjected to pressure pulses of 0.14 MPa or greater. A discussion on the experimental procedure, construction techniques, and the explosion and air-leakage test results is presented. Four types of construction material were used during this seal testing program: solid-concrete-blocks, low-density foam blocks, cementitious foam, and wood blocks. Only one of the solid-concrete block designs using a mortared thick wall survived the explosion and air-leakage tests. Most of the low-density foam block and cementitious foam seal designs were judged to meet the requirements to be considered a coal mine seal. One of the wood block convergence seal designs failed explosion testing. Based, in part, on these research findings, several alternative seal design constructions using low-density foam blocks, cementitious foam, or wood have been approved by MSHA for use in coal mines.

Jan 1, 1993

By A. J. Richardson

Before the development of the underground stoping and mining costs can be considered, certain facts about the ore body, the proposed mine, markets, etc., must be known or determined. In the case to be studied, the zinc-lead mineralization occurred with a narrow vertically dipping structure of undetermined length and vertical extent. Exploration completed to date has revealed 6.5 mil¬lion st t of proven reserves. A further 820,000 st of in¬dicated reserves has been outlined and this tonnage is considered capable of being expanded by a factor of approximately four after more detailed drilling. After studying the market conditions and completing a very preliminary feasibility study, it was decided that production would be 730,000 stpy (or 2000 stpd) of ore. First year production would be at the rate of 1500 stpd. The main design criteria for the selection of the min¬ing methods are minimizing surface subsidence, maxi¬mum recovery of the ore body, maximum degree of grade control, maximum productivity, and safe working conditions. Two basic extraction systems are considered capable of meeting these requirements: mechanized cut¬and-fill stoping and sublevel long-hole stoping with filling. The primary development system of the mine has been designed to give maximum flexibility in stoping systems and layout and to permit changes if considered necessary as a consequence of actual production ex¬perience. At the present time, access to the mine is by a circu¬lar concrete lined vertical shaft, 16 ft diam, sunk to a depth of 1380 ft. Two exploration levels have been driven within the ore zone at depths of 165 and 1246 ft below the surface outcrop. The development to date had the objective of sampling the mineralization and produc¬ing detailed information on the outline of the ore body and the distribution and controls of zinc and lead values. In an attempt to satisfy the basic design criteria for the mine, it was decided that production would be best achieved by a combination of 40% sublevel long-hole stoping and 60% cut-and-fill mining. Costs of exploration and capital development of per¬manent underground facilities are normally written off over the life of a mine. Production expenditures, on the other hand, are of a temporary nature and are normally charged as and when incurred as an operating expense. Reasonably accurate predictions of mine production costs can be built up from engineering design and estimates of individual mine activities for ultimate inclusion in the comprehen¬sive data required for financial decision making. The simulated operations can be costed on a detailed basis in the form of a monthly operating budget. The budget format can be generalized or detailed, depending upon the scope of the project. However, ex¬perience suggests that a fairly detailed format has the advantage of assuring that all significant cost items are included. For underground costing it is suggested that the budget structure include five major cost centers (i.e. development, diamond drilling, ore extraction, hoisting/ transportation, and general mine expense). These in turn are detailed under numerous subheadings. The mechanism for compiling an operating budget will be illustrated. Because of its relative simplicity, ore extraction under sublevel long-hole stoping has been chosen for illustration. All other activities, simple or complex, can be estimated in similar fashion. BLOCK AND STOPE DEVELOPMENT Long-hole blocks, used where advantageous, will be up to 250 ft in height, depending upon the vertical con¬tinuity of the mineralization, and approximately 300 ft long. Drawpoints will be at 36-ft intervals and serviced by loading crosscuts driven from a footwall drift parallel to and close to the ore zone. Pillars between the stopes will be 50 ft wide. Stopes will be drilled off with vertical rings of blastholes drilled from sublevels approximately 60 ft apart vertically. This drilling will be done by percussion drilling machines (31/2 in.) mounted on a trackless drilling rig. Load¬haul-dump (LHD) equipment will be used to move broken ore from the drawpoints to the orepass connecting to rail haulage systems. On completion, long-hole stopes will be backfilled to prevent caving and to facili¬tate later pillar removal. From a planned stope layout, a forecast of produc¬tion and development is made in Table 1. Table 1. Block Tonnage and Stope Development Quantity Ore Waste Total ore block 375,000 st 2 stopes 310,000 st 1 pillar 65,000 st Access crosscuts, 4 at 100 ft 400 ft Drill sublevel drifts, 6 at 300 ft 1800 ft Stope raises, 3 at 250 ft 750 ft Undercut sublevel drifts, 2 at 300 ft 600 ft Loadout crosscuts at 35-ft intervals 550 ft 100 ft 3300 ft 500 ft Total development footage 3800 ft Tons per ft of development 987 st

Jan 1, 1982

By Sukumar Bandopadhyay

A present dominant problem is that of high downtime of the longwall shearer used in the coal production process. The means by which this problem can be overcome are many, but one of, the most important ones is that the-use of an expert system can reduce diagnostic and repair time. In this paper, a fault-diagnostic expert system model is presented. Specifically, the diagnostic architecture of the longwall shearer expert system and its use for diagnostic reasoning are discussed.

Jan 1, 1989

By William S. Eaton

Managing a mine or mineral processing operation in today's environment presents many unique challenges to mining companies. In recent years, headlines around the world have publicized major incidents such as those associated with the Marcopper Mine in the Philippines, the cyanide release in Guyana from the Omai gold mine and numerous other mishaps associated with mining. These notable upsets were due to engineered system failures, either during normal operations or as a result of uncontrollable events. There fore, many mining firms are attempting to assess their overall risks and establish some priority for expenditure of limited operational and environmental budgets. Failure of an engineered component can have dire operational consequences that affect profitability until repair or replacement can be completed. In addition, there may be significant environmental consequences associated with the engineered system failure. This, too, may affect profitability for an extended time or even the viability of an operation. An engineered risk assessment (ERA) can support a mining company's efforts to cost effectively manage environmental risks associated with the operation of its facilities. The ERA approach also provides information about the current condition of engineered systems. This ERA approach provides a systematic review of the risks of an engineered system failure and the possible associated environmental consequences at mining company facilities. This approach has been successfully applied to many industries and types of engineered systems (Barneich et al, 1996; Whitman, 1984; Power and Tomkins, 1974; Keeney, Nair, and Kulkarni, 1979). This approach evaluates all mining and mineral processing facilities located at a single site and compares them with similar facilities located at multiple mine sites. The ERA approach has been applied to a mining, milling, smelting, leaching and refining operations. These assessments can be part of a comprehensive environmental program to aid in the recognition and cost effective management of the potential environmental risks associated with the operation of facilities. ERA results support plan-

Jan 1, 1997

By James C. Norman

The boron review arrived too late to appear in the June issue. The US was the world's second largest producer and the largest consumer of boron compounds during 1990. Boron mineral concentrate and chemical production in the US totaled 1.13 Mt (1.2 million st), including 565 kt (662,800 st) of boric oxide (B2O3) in 1990, according to US Bureau of Mines estimates. Total value amounted to $404 million. This was slightly higher in production than 1989 but down in value, reflecting a softening in boron product prices. Total world production of boron mineral concentrate and chemical tonnage in 1990 was estimated at 3 Mt (3.3 million st), which corresponds to about 1.3 Mt (1.3 million st) of BO,. World production in 1990 was about the same as in 1989.

Jan 1, 1991

By A. Marabini, G. Belardi, N. Sheau, F. Veglio

The present paper deals with the modelling of a Multi-gravity separator in the treatment of chromite fines. A first screening study was carried out by 2" full-factorial experiments in order to check the main effects and interactions among the investigated factors: shake frequency, shake amplitude, tilt angle and washwater flowrate were taken in consideration in the study. Two size fractions were employed in the experimental tests. Successively, considering the factor with larger influence on the Cr203 and SiO2 grade and Cr203 recovery in the concentrate, a further improvement of the experimental design was performed to determine empirical models to describe the behaviour of the MGS in the mineral treatment. The results show how intrinsic interaction among the factors take place, highlighting that an optimization study must be carried out considering all the MGS setting parameters.

Jan 1, 1995

By J. B. Gwiazda

Maintenance of a constant load setting throughout longwall support units and selection of the proper initial bearing capacity depending on the type of roof strata are the basic factors that ensure good performance of roof support in a longwall. These requirements can only be met by hydraulic support. The greatest advantage of hydraulic support is achieved when uniform pressure is imposed on the roof throughout the length of the longwall. Such support, however, is provided only if each of the support units acts with the same force against the roof, i.e., has the same setting load. In such cases, the roof behaves like a uniform plate without bending and shearing stresses, thereby ensuring an undisturbed structure. Without a positive set value system, achieving an equal setting load for all units of the longwall support is impossible. Due to alterations of the feed line pressure of support units as well as some reasons related to man's psychology, operators extend the height at different setting loads. This produces nonuniform roof stress, and disturbs the structure. Consequently, the roof usually cracks. The author has developed a positive set value system, which is described in this Technical Note. Selection of the setting load Two pressure values in the feed lines are usually applied in longwall hydraulic systems. Lightweight support is fed by 25 MPa (250 bar) liquid, while heavy duty units receive a nominal pressure of 31.5 MPa (315 bar). Such pressure is required not only for the props but also to power the adjust¬ment jacks and the advancing ram. If the feed pressure is too low, there will be difficulty in shifting the unit despite the inversion system of the advancing rams. On the other hand, for many roof types, the feed pressure often appears to be too high when applied as the setting load pressure. An excessive setting load acts too strongly against the roof, crushing weak strata close to the roof. The author has recognized a case where an excessive setting load destroyed not only the nearby roof strata but also the strata above a 2-m (6.6-ft) sandstone layer. In addition, an excessive setting load relieves the side¬walls, increasing the resistance when using cutting machines. As a result, the yield of coarse coal is diminished, and increased fines dominate in the final product, lowering its economic value. As indicated, selection of the proper setting load, depend¬ing on the mining and geological conditions of the extracted seam, is extremely important. In some mines, measures applied to prevent disturbances include reduction of the feed line pressure by adjusting the feed pump valve. The disadvantage accompanying reduced feed line pressure is more difficult operation in advancing the ram. Due to the reduced feed line pressure, the force of the advancing ram is much lower than the designed value. Other designs suggest using a third feed line. However, installation of supplementary valves on the support units is required, a time-consuming and expensive procedure. The disadvantages of the powered supports are eliminated by a system designed by the author. So far, such a method of setting load control has not been used in any type of support. Setting load control unit The designed positive valve set for prop loading and the setting load control correspond to existing control systems for hydraulic powered support. The layout of the unit connected to the hydraulic prop control is presented in Fig. 1. The unit is marked LIDS. It incorporates three valves that may operate separately or connected. Valve A automatically opens and closes with liquid flow in the prop feed circuit. The valve is opened when the canopy touches the roof and closed when the support unit is withdrawn. Valve B serves as the setting load control. Valve C automatically opens and closes the flow in the line connecting the under-piston space of E to the prop F with the separator G. The valve block of each support prop is marked BZ. The UDS unit is connected by the hydraulic lines to the F prop control circuit. A valve is connected by H to pressure line J and by K to the G separator. In the UDS-3 version, line L is connected to M, linking the over-piston space of F prop with the G separator. Valve B is linked with valve A by a connector; it is also connected to the under-piston space E of prop F by line P. Valve C is fixed between lines K and P, connecting space E of prop F with the separator G. When setting the support, the liquid flows from line J through separator G, the BZ valve, and valve C to the space E of prop F. When reaching the roof with the canopy, valve A is opened and C closes. In this way it is impossible for the operator to cause the liquid pressure in space E of prop F to reach the level of line J. The prop pressure is set by valve B of the UDS unit. When withdrawing the support, valve C is automatically opened and A closed. Three UDS units have been fabricated and are designated

Jan 1, 1990

By Pradip

The selection, testing, and commercial utilization of flotation reagents is best carried out in stages through microflotation tests on pure minerals and their synthetic mixtures followed by bench-scale batch flotation tests on the ore. A systematic approach to quantitatively evaluate different reagents, in particular, through the analysis of batch flotation data, has been presented in this paper. It is recommended that the reagents are compared both in terms of their performance with respect to the recovery-time as well as the recovery-grade plots. While the former describes the kinetic aspects of reagent interaction, the latter is a measure of its selectivity for a particular separation problem.

Jan 1, 1987

By R. L. Grayson

In 1995, the United States Congress abolished the Bureau of Mines in the US Department of Interior. But Congress preserved the safety and health research functions being performed at the Pittsburgh and Spokane Research Centers. These functions were transitionally housed in the US Department of Energy before permanently merging with the National Institute for Occupational Safety and Health (NIOSH), Department of Health and Human Services, in October 1996. Since the rnerger of the former Bureau of Mines with NIOSH, there has been an active evaluation of ongoing research activities, significant interaction with stakeholders, and the development of a process leading to "data-informed" priority setting. This paper reviews the current surveillance information concerning the level of injury and disease from mining work and discusses how research priorities are evolving to improve the contribution of research to the prevention of injury and disease.

Jan 1, 2000

By Raj K. Rajamani, M. K. Abd El-Rehman

The planetary mill is very much suited for the ultra-fine grinding of mineral fillers. Ultra-fine grinding here refers to grinding powders to a size range 5-1 0 rnicrons. Such powders enhance the quality of the final product. The operation of a specially built planetary mill is described. It is shown that a target size of 5 rnicrons can be reached in a few hours using the dual drive planetary mill. A centrifugal field of 37G is generated by the gyration arm of the mill. The design allows the desired setting of canister critical speed. The ultra-fine grinding of alumina, titanium dioxide, hematite, magnetite, kaolin and silicon carbide are described. Also, the continuous operation of the mill is shown.

Jan 1, 2003

By Photios G. Ioannou, Guillermo F. Salazar, Herbert H. Einstein, Yeong W. Kim

The MIT tunneling group has developed, over the past decade, a number of decision support tools for the geologist and engineer, planner and designer, owner and contractor involved in underground construction. These computer based tools can be used in the planning and exploration stage of a project, in predesign and design, in construction planning and during construction. Tasks as different as optimizing information collection during exploration and during construction, as selecting major design alternatives in predesign and modifying design during construction, as providing cost and time estimation and considering social and physical environmental constraints, can be handled. Output can be in the form desired by the user ranging from geologic profiles (cross sections) to management reports.

Jan 1, 1987

By E. S. Peterson, M. K. Harrup, K. A. Dolbeare, J. E. Wey

Efficient methods for the removal of copper from heap-leach solutions without mobile organic phases are a current mining industry need. To satisfy this need, a selective, high-capacity, durable solid material is required. The Idaho National Engineering and Environmental Laboratory (INEEL) developed a set of materials based on current solvent-extraction chelates that provide selective copper concentration in the presence of high concentrations of iron. In contrast to conventional solvent extractions, the author's strategy anchors the chelators to a nonpolymeric solid substrate through the use of a custom-designed linker molecule. The fundamental synthetic routes to these materials have already been established, and attachment to the solid surfaces have been achieved in an efficient manner. The author's continuing research is focused on the optimization of the synthetic pathways to these solid materials and the evaluation of their copper-capturing capabilities.

Jan 1, 2003

By P. E. Nielsen, T. L. Loga, B. B. Mauko

A numerical model is presented which simulates the dynamic fracturing process for a circular geometry fracture in a uniform stress medium. The dependence of crack aperture and net pressure in the crack upon various dominant parameters of the fracturing process is discussed. An analysis of a 75% nitrogen foam fracture in a coal seam at 5600 feet (1707 m) is also presented. The analysis is based upon the discussed parametric trends, the job records of treatment parameters, and a Nolte plot of net downhole pressure. It is interpreted that fracturing with water, in pre-foam fracture pump tests, resulted in significant fracture height growth. However, the subsequent foam fracture had reduced height growth due to the high effective viscosity of the foam and due to the higher magnitudes of elastic modulus and stress in the bounding rock strata with respect to the coal.

Jan 1, 1986

By Paul H. Lu

This paper proposes a novel method, "integrity factor approach," of mine-pillar stability evaluation based on in situ triaxial-loading measurement. The pillar integrity factor is defined as the quotient resulting from division of integrated total pillar strength across the critical dimension by the integrated total pillar load across the same dimension. The integrity factor is dynamic, while the conventional safety factor is static. With this approach, the pillar stability can be evaluated at any stage of mining progress during panel operation. It automatically considers the progressive failure of the pillar and post-failure strength without the need to clearly define the boundary between yielded and intact zones of the pillar. Based on two case studies, the integrity factor approach is shown to be a practical, effective technique for evaluating stability of conventionally designed pillars. It is suggested that a more realistic estimate of in situ pillar strength could be made based on the combination of in situ triaxial-loading measurements and laboratory true triaxial compression tests, rather than through the traditional combination of biaxial-loading measurements and conventional triaxial compression tests.

Jan 1, 1986

By Jacques L. Mauboussin

The EUROPEAN ORGANIZATION for NUCLEAR RESEARCH (CERN) has undertaken the construction of a LARGE ELECTRON POSITRON COLLIDER (LEP), near GENEVA (Switzerland). The facility includes the main 27 km long tunnelled ring, 18 shafts, 20 experimental caverns and many auxiliary galleries. Roughly 800 000 cu.m were excavated in the molasse at 50-140 m depths, using roadheaders for caverns and auxiliary tunnels and 3 WIRTH, 4.5 m diameter, fullface TBMs for the main circular tunnel. Each TBM performs 25-30 m/day. Provisory lining behind them uses either precast concrete elements or shotcrete. Works a e being carried out by a joint-venture of European contractors led by FOUGEROLLE (France).

Jan 1, 1987

By R. Pakalnis

This paper compiles the research work conducted by the Spokane Research Laboratories in the investigation of shotcrete within underground mines to the applied design of ground support for the operator. It looks at the design methodology in terms of employing shotcrete as a fabric to confine the rock mass between the individual bolts as dictated by empirical charts. The failure mechanism in terms of its implication on design is presented with respect to the design of 6m wide tunnels for temporary and permanent openings within weak rock masses. This is reflected by field observations that relate laboratory tests to the onset of failure for the operator in order to provide guidelines for design and monitoring. This is coupled with guidelines for safe entry in terms of cure times as well as the understanding of the behaviour of the shotcrete in terms of fibres and the overall strengths that result with respect to the design of the mine openings. This enables the operator to better understand his workplace thereby provide a safer work environment.

Jan 1, 2010

By R. Hogg, A. C. Maffei, D. T. Ray

Based on standard population balance models for agglomeration and size reduction processes, a simplified model has been developed to describe simultaneous growth and breakage in the flocculation of agitated suspensions. The results of semi-batch flocculation tests have been used for estimation of the model parameters and empirical relationships have been obtained to describe the effects of process variables and to establish scale-up criteria. Application of the model to the design, operation and control of continuous flocculation processes is discussed.

Jan 1, 1990