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By D. Burkhard, E. Herth

"Geologic structure and properties are the primary influences on ground control at San Juan Mine. Geotechnical properties of the coal seam and near seam strata, the presence of above seam sandstone paleochannels and water, are thought to be the leading contributors to ground control issues. It is the confluence of specific geologic conditions rather than isolated features that most often produce adverse conditions. Altering primary support installation and mining practices are done when unfavorable conditions appear during development. Installation of standing support is most often employed when adverse conditions develop at later dates. REGIONAL GEOLOGIC SETTING San Juan Mine is an underground coal mine located on the western flank of the San Juan Basin in northwest New Mexico. The San Juan Basin is contained within the south eastern region of the Colorado Plateau geographic province. The basin contains multiple coal bearing formations and is also a major producer of coalbed methane gas. San Juan Mine is stratigraphically located within the Fruitland Formation. From a prominent monocline marking the western boundary of the basin, the formation dips gently to the east towards the center of the basin. The primary driver for the deposition of coal bearing formations within the basin was the Cretaceous Inland Seaway. Multiple cycles of transgression and regression driven by sea level rise and tectonic subsidence deposited rock types from marine shale’s to terrestrial sandstones. The Sevier Mountains to the west of the seaway likely provided the sediment supply for the Fruitland Formation. During a major cycle of seaward regression, the Pictured Cliffs Sandstone Formation prograded over the marine Lewis Shale. The coal bearing Fruitland Formation followed, being deposited over the Pictured Cliffs. The Fruitland Formation contains numerous coal seams and is approximately 400 ft. thick at San Juan Mine. Fruitland Formation coal seams are numbered in order of superposition. San Juan mines the Fruitland number 8 seam. Dating has placed the age of the seam at approximately 73 to 75 million years [1]. LOCAL GEOLOGIC SETTING The Fruitland 8 coal seam dips approximately 1-3 degrees towards the E-NE. At San Juan Mine, the seam displays a mild plunging, synclinal structure, the axis roughly bisecting the mine from west to east. The number 8 seam consists of five coal benches and four volcanic ash band partings labeled sequentially based on superposition. The lowest portion of the seam is the 8-1 bench, the 8-2 through 8-5 benches following sequentially to the top of the minable seam. Ash band partings are labelled P1 through P4 from bottom to top respectively. The presence of distinctive partings and benches allows for seam definition and aids in identifying mining horizons."

Jan 1, 2016

By O. Lopez

"Stringent regulations and increased arsenic presence in ore bodies being exploited have created the need for cost effective and environmentally compliant technologies to remove selenium and arsenic from mine waters and metallurgical waste streams. Existing treatment methods have high life cycle costs and/or produce residues that, due to their unknown stability over the long-term, may represent future liabilities. This results in high compliance costs and uncertainties which negatively impact existing operations and projects in permitting. Selen-IX™ is a new ion exchange based process that selectively removes selenium from mine impacted waters down to 1 ppb while producing a stable inorganic iron-selenium product. Results of piloting completed on streams generated from coal and copper-gold deposits are presented along with preliminary equipment sizing and life cycle costs of full scale treatment. An example of a commercially available system for arsenic removal using sulphide precipitation is presented. The benefit of this treatment method is discussed based on recently completed technical feasibility assessment including laboratory testing and preliminary engineering design. INTRODUCTION Selenium Due to the chemical similarity between selenium and sulfur, selenium substitutes sulfur in many minerals that can be found in a wide range of deposits including coal, base metal, precious metal, phosphate, and uranium. One of the main environmental challenges facing the industry in North America today is related to the removal of hexavalent selenium, e.g. selenate, to ultralow levels, typically between 1 to 10 ppb, from large flows of often cold water containing less than 1,000 ppb of selenium. Recent studies commissioned by the industry identified biological selenium reduction systems as the best available technology. Although the biological selenium reduction process is commercially proven in certain applications to remove selenium to ultralow discharge levels, the technology has been primarily applied to relatively low and constant flows of warm water produced by the Flue Gas Desulfurization (FGD) scrubbers in the power generation industry. The wide spread adoption of this treatment method by the mining industry is hindered by high capital and operating costs associated with treating large flows of cold water, lack of data and knowledge about the long term stability of solid residues produced by this treatment, and limited capacity of biological systems to respond to sudden changes in flow and/or mass load of contaminants. Furthermore, it has been reported that reaching discharge limits less than 3 ppb could be a challenge and that polishing treatment for the removal of BOD and phosphate maybe required to avoid the impact of the biological treatment which requires organics and nutrients, on aquatic environment downstream of treatment plants. Alternate treatment systems are being developed to overcome and/or completely avoid some of the challenges of biological systems. One such treatment system is BioteQ’s Selen-IX™ process which is purely physico-chemical treatment system."

Jan 1, 2015

By T. Karlkvist, H. R. Kota, F. Andreas, B. I. Pålsson

"Separation of different calcium minerals have long been an interesting and challenging problem. In this investigation calcium mineral separation is examined by: microflotation, zeta potential measurement and adsorption, using novel collectors having two functional groups instead of one. In theory, by varying the distance between the functional groups, it could be possible to preferentially target one calcium mineral by matching the spatial distance between the sites on the mineral surface. In this investigation two new surfactants have been tested to estimate their ability to float apatite and/or calcite. Preliminary findings show that an increase in distance between the functional groups favors one mineral over the other, and this might be due to differences in the mineral surface structure. INTRODUCTION Apatite is a strategically important mineral which has both great value for industrial use as a precursor to phosphoric acid; and in the agricultural industry as an important source of phosphorous in fertilizers. This has led both the European Union and the United States to recognize it as an important resource for the future. Currently there is no lack of apatite in the world, but it is important to develop the beneficiation to ensure future supply. A common problem in apatite flotation is the selectivity towards calcite, which it is commonly associated with. Although calcite is an important industrial mineral its abundance makes it not usually worth to extract from apatite ores unless it is easy to separate1,2. Both apatite and calcite are semi soluble minerals with calcium as one of their constituents. The most common way to separate apatite and calcite is to use flotation. A process that involves making the mineral particles to extract from the system hydrophobic by adsorbing a collector on the mineral surface. One way to study collector adsorption on a mineral surface is by zeta potential measurements. The zeta potential is the charge at a slip plane a short distance from the surface. By measuring the zeta potential with and without the presence of collector it is possible to investigate if the collector is present at the surface or not, a change is indicative of collector adsorption. The zeta potential of fluorite, apatite and calcite without collector can be seen in Figure 1. The small difference in zeta potential between apatite and calcite shows the similarity in their surface properties and gives an indication of their separation difficulty. One additional problem with semi soluble minerals is their tendency to slightly dissolve in the pulp and thereby change the ionic composition of the solution. This makes separation and research problematic as the freed ions may readsorb on the mineral surface and interfere with the collector adsorption on the mineral surfaces3–5."

Jan 1, 2016

By L. Yan, C. Zhou, M. Reyes

"The MINER Act requires the installation of post-accident, two-way communications and electronic tracking systems for all coal mines. One communication system which can satisfy this requirement is an electrode-based, through-the-earth (TTE) communication system, which sends signal directly through the overburden of a mine. The performance of these types of systems relies on several controlling factors such as earth conductivity, overburden depth, operating frequency, TX/RX antenna length, TX/RX offset, TX/RX orientation angle and TX current. In an effort to estimate attenuation through the earth, NIOSH researchers developed an analytic solution for the electrical field distribution of an electrode-based TTE communication system for the extremely low-frequency (ELF) range (< 30 Hz) or DC excitation source. The analytic model was then validated using test data measured at an operating coal mine. The validated model can be used to optimize the operating frequency, estimate the required transmitting current, choose appropriate TX/RX orientation, and predict the performance of an electrode-based TTE system at a coal mine. INTRODUCTION One of the requirements of the Mine Improvement and New Emergency Response Act (MINER Act) of 2006 is the installation of post-accident, two-way communications and electronic tracking systems for all underground coal mines. As one of the through-the-earth (TTE) wireless communication methods available for the mining industry, the electrode-based, or linear, TTE systems establish communication by inducing a current within the earth medium by use of a pair of current probes. The resultant current is detected as a voltage by a similar set of distant probes. These probes could be an insulated piece of wire tied to two electrodes grounded at distant ends, or the grounded ends could be rails, roof bolts, or grounded electric rods. The analytical solution developed in a homogenous half-space, for this form for various transmitter/receiver (TX/RX) configurations, will be presented in this paper. The communication signals can be sent either from underground to surface (uplink) or from surface to underground (downlink). In order to achieve a successful communication link, it is necessary to understand the electric fields within the earth so that the appropriate size of transmitting antenna, TX/RX relative orientation, and operating frequency can be chosen. Hill and Wait gave the explicit forms of partial magnetic fields and partial electric fields of electrode-based TTE downlink communication with finite cable length [1]. In prior research, the E-fields of electric line transmitter were obtained for a static TX current [2] [3]. In this paper, the general and explicit forms for electrode-based TTE electric fields are developed and presented. The solutions for the extremely low-frequency (ELF) range (< 30 Hz) or DC excitation source obtained from these forms are presented. Also, factors that affect the E-field strength at RX are discussed."

Jan 1, 2017

By N. Santa, V. Salinas

"Printed circuit boards are potentially recyclable materials. Developing suitable methodologies for proper disposal and better utilization, brings major environmental and economic benefits. A methodology for copper recovery using heat treatment before conventional leaching is carried out. A comparison between conventional methodology and the one that includes incineration treatment is presented. It concludes about results and differences in terms of the percentage of copper leached over time. INTRODUCTION The rapid proliferation of electronic devices in the past two decades, has forced researchers to find a solution for one of the most toxic and hazardous materials, the printed circuit boards (Ghosh et al, 2015). For this, countless techniques have been developed, the methodologies includes hydrometallurgy, pirometallurgy, electrometallurgy or a combination between all of them. The amount of precious and base metals in its composition are very attractive and usable to implement methodologies that allow maximum recovery and recycling. This subject has been extensively studied and researched in order to provide solutions to the great problems generated by incorrect final disposal of this waste. Over 150 articles have been widely published in the last 15 years, covering large areas such as characterization of discarded printed circuit boards, health risks associated with processing and different routes of recycling have been analyzed to provide an overview understanding this issue (Ghosh et al, 2015). (Oh et al, 2003) performed a selective leaching of valuable metals from waste printed circuit boards by size reduction, magnetic and electrostatic separation, also leaching with sulfuric acid and hydrogen peroxide, achieving 95 \% of copper extraction. (Chaverra Arias, 2014) proposed a similar methodology to extract copper from electrical and electronic waste, performing similar processes and applying metallic copper electrowining as well as copper oxychloride precipitation. (Silvas et al, 2015) used hydrometallurgical processes for recycling printed circuit boards applying physical processing and copper extraction by selective leaching. (Kumar et al. 2014) conducted a leaching of metals from waste printed circuit boards using sulfuric and nitric acids. Also there have been proposed various methods for recovery of different metals including applying multiple metallurgical techniques that involve both living organisms and high temperature processes such as pyrometallurgical. The printed circuit boards are a very important source of copper, according to several authors mentioned above, the amount of copper on a printed circuit board can reach about 20 \%, because copper is responsible of establishing a conductive layer that will enable the connection between electrical and electronic components. This document proposes a methodology for recycling printed circuit boards and extract copper by selective leaching. The process that is carried out consist in a preliminary characterization and mechanical separation which will condition the material for further hydrometallurgical treatment that includes leaching in nitric acid media 3M."

Jan 1, 2017

By C. Anderson, B. Schug, C. Cristoffanini, S. Nazari

The Haile Gold Mine in South Carolina is increasing the automation level of their processing plant in order to improve operational stability. One step towards this goal is the application of a Digital Twin concept to provide Virtual Instruments. Haile has engaged ANDRITZ to implement Virtual Instruments at their grinding area hydrocyclones and at their tailing thickener underflow pump. These virtual instruments provide density values in a traditionally difficult location for reliable slurry density measurement. The Digital Twin is a first principles IDEAS1 ™ simulation model, connected in real time to the control system. Results have been promising, with the virtual instrument values matching the physical instrument under normal operating conditions, and showing better prediction under upset conditions. Haile plans to use the virtual instruments as optional process control parameters. INTRODUCTION OceanaGold’s Haile Gold Mine is located near Kershaw, South Carolina, about 50 miles southeast of Charlotte, North Carolina. This location was first mined in 1827, and has produced gold through several operating periods since then, including a heap leaching operation from 1985 through 1992. In October, 2015, OceanaGold acquired Romarco Minerals, the previous owner of the Haile Operation. The current operation began in 2017 as an open pit mine with additional underground resources. The nameplate throughput is 6,350 tpd ore, and 2017 production was 118k oz gold, with 140 to 155k oz gold expected in 2018. Proven and Probable reserves exceed 3 Moz. The process, as shown in Figure 1, includes a primary jaw crusher, comminution with a Semi-Autonomous Grinding (SAG) mill and ball mill circuit, flash flotation, rougher flotation, regrinding, pre-aeration, carbon in leach (CIL) leaching, carbon elution and regeneration, cyanide recovery, electrowinning and refining. The Haile Operation’s inaugural gold pour took place January 20, 2017. Mine life is expected through at least the year 2030. Based partially on prior experience of some key OceanaGold personnel with automation services from ANDRITZ on other projects, OceanaGold decided to include ANDRITZ on the Haile project. The scope of these services for Haile included design and supply of the process control application software, an ANDRITZ dynamic process simulator for operator training, and ANDRITZ Advanced Control Expert (ACE) ®2 systems for the SAG mill, ball mill, and rougher flotation areas. Since startup, ANDRITZ has continued to provide services for the site, particularly related to PLC and HMI equipment, programming and support, ACE system expansion, and the implementation of Virtual Instrumentation. The implementation of this new Virtual Instrumentation technology at Haile is consistent with Haile goals: increased automation, a change in the role of the operator from interactor to monitor, and improved stability by removing human error.

Jan 1, 2019

By John W. Rushton

Maintenance and inventory are usually the last items considered in mine development. They need to be addressed properly and early. The micro-computer brings sophistication within the reach of even the smallest mines. Maintenance and inventory programs, if properly installed, will reduce costs 15 to 25% and allow management time to address the mining and metallurgical problems at start up. Shops and warehouses are rarely designed properly. Large engineering firms spend 5 to 20 times more than is necessary for first class facilities. There is an optimum design, size, and layout. Take the time necessary to assure you get the right people, programs, and facilities in that order. Maintenance, inventory, and inventory control must be addressed early if a small mine is to have a chance at optimizing costs during the first two years of operation. This is hard work and is usually put on the back burner until the mine is in some degree of trouble. Hiring an experienced maintenance superintendent is important but it is not the complete solution. Maintenance systems, design of facilities, and use of computers are not skills that maintenance people usually have. In addition, the work of ordering spare parts and working with erection people is a full time job and any maintenance people on site during construction are quickly recruited. The end result is that few mines start up with an adequate Maintenance or Inventory System in place. The months, and in some cases years, that the mine attempts to operate without a system are both painful and expensive. Designs of successful integrated maintenance and inventory programs are virtually 100% supported with some outside assistance. Our system is the approach that we are most familiar with and this paper will isolate how Rushton International would go about setting up systems and designing support facilities. Micro-computers are the only way for a small mine to go. ~ini-computers cost more originally and operating costs are many times higher than even the most sophisticated micro system. Micro- computer software is developed further and recent advances in hardware makes medium priced units equal to most mini&apos;s in RAW COMPUTER POWER. Mines in general, and especially small mines, have no need for main frame power. MAINTENANCE A good maintenance program has: 1. Equipment Master List a. Equipment Detail Reports 2. Work Order and Backlog Control Program 3. Component Control Program 4. Tire Program 5. Wear Materials Program 6. PM Control Program 7. Maintenance History Program 8. Performance Program We also track miscellaneous items such as MSHA training and inspections, timekeeping, etc. Rather than going through the program step-by-step, it makes more sense to go through a typical Maintenance Day and show how the programs are used. The first thing maintenance supervision needs to know is "how we are doing." We

Jan 1, 1987

By Nick Chlumecky

INTRODUCTION Column-type supports are used to prevent roof falls in underground excavations. The first individual column supports were made from timber that was cut to fit the irregular heights of mine seams. Because the seam height of the material removed varied throughout the mine, timber posts generally could be used only once. During the I 920s, steel friction props began to appear in British and European mines, and, by the 1950s, thousands of these props were being used. However, the steel friction props were heavy, difficult to install, and difficult to recover. Furthermore, the friction props could not be preloaded, and their yielding behavior was rather erratic. In 1946, hydraulic props were introduced in En¬gland, and, within a few years, they replaced friction props throughout Europe. The first application of hy¬draulic props in the United States occurred in the 1950s, during early experiments with longwall mining. Modern technology and high-production mining procedures have combined to make the use of hydraulic props a very acceptable method of providing temporary heavy-duty support for production areas. In areas such as development headings, where only modest roof loads are expected, lighter-duty jackscrews, ratchet jacks, and light hydraulic props are used. ADVANTAGES OF HYDRAULIC SUPPORTS The acceptance of hydraulic supports has been wide¬spread during the past 25 years, proving their reliability and effectiveness in preventing roof falls. Some of the advantages provided by hydraulic supports are: 1) Individual hydraulic props can be installed within 30 sec, resulting in a minimum exposure of per¬sonnel and equipment to an unsupported roof. 2) Hydraulic props can be preloaded to more than 4.5 t (5.0 st), increasing the frictional resistance between the rock strata. Thus, the roof is kept tight, improving the behavior of the unsupported ground. 3) If gradual roof sagging or bottom heaving im¬poses an irresistible force, the hydraulic prop yields gradually and continues to support the ground. 4) The support capacities of the hydraulic props most widely used range between 10 and 20 t (11 and 22 st). Their support capacities are maintained during yielding to an irresistible force. 5) Hydraulic props can be recovered rapidly. Where no other supports are provided for the roof, as in pillar¬ing operations, the hydraulic props can be recovered by remote means. 6) The length of hydraulic props is readily adjust¬able allowing them to be reused in seams where roof height varies. 7) Studies of mining operations indicate that the use of hydraulic props increases both personnel safety and mine productivity. DESCRIPTION Figs. I and 2 illustrate typical hydraulic props used in mines in the United States. Fig. I illustrates a prop

Jan 1, 1982

By Gary Krix

The Amdel Thickener Interface Gauge (TIG) provides continuous on-line measurement of the level of the liquid/solid interface in thickeners and Counter Current Decantation (CCD) washers. Unlike other interface level measurement systems, the TIG also determines the span of the transition zone. The transition zone is the region between the clear liquid overflow zone and the compressed solids zone. The Thickener Interface Gauge determines the level of the interface by measuring the natural radioactivity of the solids in the thickener tank using an array of up to twelve detectors mounted vertically in a stainless steel probe. All solids contain low levels of natural radioactivity while the liquid phase generally does not. The concentration of natural radioactivity at each level in the thickener therefore depends on the concentration of solids. The radiation emitted by the solids is converted to electrical pulses by the detectors in the probe and the pulse-rate provides an accurate measurement of the density. A typical count-rate profile is presented in Figure 1. The vertical axis indicates the depth in the tank referenced to the surface of the liquid. The horizontal axis indicates the count-rate measured by the TIG probe, this count-rate is proportional to the density and can be used to determine the settling rate. The count-rate at the top of the graph is relatively low, this region corresponds to the clear overflow zone which contains virtually no solids. Moving down the vertical axis the count-rate increases dramatically at the level of the transition zone where the solids content increases with the depth in the tank. In this region the settling conditions of the solid particles change from free settling to compression. Moving further down the vertical axis the count-rate reaches a plateau where the count-rate remains relatively constant although there are often small fluctuations, this region corresponds to the solids compression zone. The liquid/solid interface as measured by the TIG corresponds to the depth at which the count-rate crosses the target count-rate which is generally selected to be midway between the liquid count-rate and the compressed solids count-rate. The length of the TIG probe is selected to cover the typical operating range of the liquid/solid interface in the thickener or washer. If, for example, the minimum bed level is 3.0 metres below the surface of the liquid then the probe length is selected to extend 30 metres below the surface. The TIG probe can include up to twelve detectors and the se are spaced over the measurement range. For small measurement ranges of around one metre 4 to 6 detectors would normally be used while for larger measurement ranges of three metres or more twelve detectors would be supplied.

Jan 1, 1993

By T. W. West

INTRODUCTION This paper presents an assessment of conventional flotation technology. In addition it presents some ideas which can be used to change the perception of what may be done with conventional flotation equipment to maximize its efficiency. Perspective of Conventional Flotation Technology 1. Flotation in many coal preparation plants is defined as something between a necessary evil and a non-viable concept because: -Too much art is required on a continuous basis to achieve uniformly acceptable results. -It is expensive, the capital cost in dollars per ton of circuit feed for flotation and dewatering is high. -It is a difficult process to use when SO2 reduction governs system design and equipment selection. -It creates a product (fine coal froth) that is difficult to dewater in the presence of BTU/lb as-received pressures. -It is a difficult process to control from a management standpoint, there are no dependable set points to rely on. 2. Many froth cell circuits are 0.5m x 0 (nominal 28 M x 0) feed which is sourced from screen/cyclone underflows. The original intent was to obtain maximum yield at acceptable product quality and moisture using disc filters. (See Figure #1) 3. An alternate to conventional flotation and its perceived to be limits of efficiency, capital costs, etc. is to move towards such developing flotation technologies as column flotation, reverse flotation, and microbubbles. A lot of work is being done to move these concepts thru pilot stages into full scale, commercial plant operation. The majority of the coal producers and coal preparation plant designers view these concepts as being in the development/pilot plant stages. OBJECTIVES Based on the preceeding comments of the perspectives of the coal producers and plant designers, the options are limited. They can make do with what they have available to them in the way of industry standard equipment and practices or they can move forward with developing technology. I feel there is a third option: what can we really do today with what is available to us, without reinventing the wheel? APPROACH 1. When processing coal against ash and SO2 restrictions, the fine clean coal stream can be the best stream on the clean coal conveyor due to the release by size and the ability to create a high BTU/low sulfur (dry basis) product. 2. Targets: -Consistent raw coal size consist in the plant feed . -Do not put pyrites in the froth cells. -Adjust criteria defining good flotation away from yield towards maximum BTU recovery at minimum ash. -Limit the froth cell feed size consist to 0.15 mm x 0 (nominal 100M x 0). -Establish and maintain a correct, consistent percent solids in the froth cell feed. -Select flotation reagents with care. -Introduce reagents in the proper sequence.

Jan 1, 1988

By Neil J. Dahmen, Howard J. Handewith

INTRODUCTION Tunneling machines were introduced into civil con¬struction in the 1800s, but their practical applications were not realized fully until the 1950s. As of 1977, tunneling machines were being used on most tunneling projects of substantial size. As in civil construction, the mining applications of tunneling machines have included long drives such as tunneling for long ore-body develop¬ment drifts, long haulage tunnels, longwall access drifts, and long crosscut developments. Tunneling machines ranging from 1.75-m (5.7-ft) to more than I1.0-m (36-ft) diam are available from several manufacturers; machines for mining applications generally have ranged between 1.75 and 6.0-m (5.7¬and 20.0-ft) diam. Tunneling machines usually operate at or near horizontal grades and are supported by a rail-haulage backup system. However, declines and inclines have been made using alternative backup equip¬ment. Either circular or shaped tunnels may be bored, depending upon the particular equipment used. Using tunneling machines to bore openings through rock offers several advantages over the use of the drilling and blasting technique. These advantages include a rapid advance, continuous operation, uniformity of the opening size, smooth walls, and safe working conditions. Compared to drilling and blasting, boring reduces the disturbance of adjacent rock, thereby reducing the ground-support requirements and the exposure to water inflows. The smooth walls provide natural arching and present minimum resistance to the flow of air. A tunneling machine should be considered a spe¬cialized mining tool that is well-suited to certain mining applications. Properly applied and sustained, a tunnel¬ing machine contributes to realizing substantial econo¬mies in the development of underground mines. All costs herein are expressed in 1977 US dollars, unless otherwise noted. The mention of trade names and manufacturers is not intended to imply an endorse¬ment of any specific equipment. TYPES OF TUNNELING MACHINES As listed in Table 1, five principal types of tunneling machines were being manufactured in 1977. Each type of machine offers a special solution to a specific mining problem. Although each of these types of machines is described herein, most of this chapter is devoted to the full-face rotary rock-tunneling machine that has demon¬strated the greatest production in both civil construction and mining applications. Full-Face Rotary Rock-Tunneling Machines As shown in Fig. 1, a full-face rotary rock-tunneling machine consists of a rotating head that is fitted with rock-cutting tools and is forced into the tunnel face. This head is supported by a bearing on a structural support member that, in turn, is held in place by a hydraulically positioned wall-gripping mechanism. Both torque and thrust are reacted from the cutterhead, through the structural support and gripper mechanism, and to the tunnel wall. The rock cuttings fall to the invert at the tunnel face, and they are removed by means of cutterhead buckets or scoops that transfer the cuttings to a conveyor belt. After advancing the cutter¬head through a preset boring stroke, the tunneling machine is advanced by hydraulically pulling the gripper mechanism in from the tunnel walls, stroking forward, and resetting the gripper to the new forward position on the wall; the machine then is set for the next advance stroke. Full-face machines cut only circular openings. Ma¬chines with diameters as small as 1.8 m (6 ft) and as large as 6.1 m (20 ft) have been used in mining applica¬tions, excavating rock with unconfined compressive strengths in excess of 276 000 kPa (40,000 psi). Although special machines have been designed to turn on a radius of 30 m (100 ft) in both horizontal and vertical directions, the conveyors and backup sup¬port equipment generally limit tunneling machines to turns with a radius of more than 60 m (200 ft). The full-face rotary rock-tunneling machines have provided the fastest and most reliable production of any mining method or machine. These machines can be equipped to place any of a variety of ground-support systems, and they can be provided as "open" machines, as shown in Fig. 1, or as "shielded" machines, as shown in Fig. 2. The open machines have the advantage of allowing the ground to be supported as near to the tunnel face as possible. The shielded machines have been designed principally to allow placing precast con¬crete segments. The shielded machines completely pro¬tect the equipment and personnel at the heading, but they are economically limited to permanent long-life haulageways where full lining is required. Partial-Face Rotary Rock-Tunneling Machines As shown in Fig. 3, the partial-face rotary tunneling machine is equipped with a cutterhead that rotates slowly at about 0.21 rad/s (2.0 rpm). The single-disk cutterhead produces the tunnel section shown in Fig. 4 by means of an oscillating motion. These machines work on the undercutting principle, and they are avail¬able in spring-line dimensional increments of 0.1 m (0.32 ft) from 1.5 to 2.1 m (5.0 to 7.0 ft); the vertical dimensions are 2.4 to 3.2 m (8.0 to 10.5 ft), respec¬tively. Depending upon the tunnel width, the cutterhead disk may be fitted with from 8 to 16 drag-bit cutters. Penetration is achieved by means of hydraulic rams that force the tungsten-carbide bits into the rock mass. The operation is cyclic, with cutting occurring only on the upstroke. On the downstroke, the rotating cutter¬head draws the fallen muck from the invert into the panzer-type conveyor for rearward transfer. This feature allows time to either reposition ore cars or to exit the tunnel with muck trains while the machine is cutting, as shown in Fig. 5. Boom-Type Rotary Excavators The boom-type rotary-excavator tunneling machines commonly are known as boom-type continuous miners

Jan 1, 1982

By B. A. Ferguson, P. W. MacMillan

INTRODUCTION Jack Spalding in Deep Mining, Chapter 3, states: "In deep mining, to start stoping an orebody without a definite plan of operations covering the whole extraction from beginning to end is to invite serious trouble. Any stoping without plan is bound to leave, toward the end, a number of pillars, remnants, or promontories, which, as they are reduced in size, are liable to fail and cause a general collapse. The object of stope planning is so to mine an orebody that at no stage in the operation is a remnant left. Definition: `When a block of ore is stoped in such a way that eventually a small piece is left entirely surrounded by stoping, that piece is termed an island remnant.&apos; The design of and adherence to a plan of stoping is of greater importance in preventing rock bursts than the type of ground control adopted." This, of course, is just an example of the old axiom that an ounce of prevention is worth a pound of cure. And in this context, mining does not have to be very deep before poor ground conditions, made worse by lack of good planning, force a mining company to resort to more expensive methods of ore recovery. At all times when rock pressures become excessive, stoping plans must be devised to avoid the creation of pillars or promontories. This is most readily recognized when mining deep narrow ore veins by open methods. It would be a little more general to say that stoping systems should seek to avoid or minimize localized concentrations of rock stress. Ore bodies, and particularly those at Falcon¬bridge, can reach large widths and leaving pillars for extraction at a later time cannot be avoided. However, good planning can reduce the effect of factors which create bad ground conditions. The necessary ingredients of planning are time and information: time to formulate and revise ideas; information, gained in exploration and early development, to provide the fullest knowledge of the ore body. In this way, plans become meaningful. It must always be kept in mind that the long-term results of good planning often require decisions not compatible with reaching earliest full production. PLANNING FOR DEEP MINING At Falconbridge Nickel Mines ore bodies presently being found and prepared for production lie at depths in excess of 610 m (2000 ft) below surface. Current un¬derground exploration is being directed towards favor¬able locations below 914.4 and 1219.2-m (3000 and 4000-ft) depths, at most mines. The Onaping Deep ore body has been traced to a depth of 1508.8 m (4950 ft) and at Falconbridge mine the ore body is under develop¬ment below the 5000 level. Planning for more efficient ground control has therefore become an important con¬sideration of future mining. HYDRAULIC BACK FILLING AND CUT-AND-FILL STOPING The advent of hydraulic back filling has led to the increasing use of flat-back cut-and-fill stoping methods during the last ten years, until it has probably become the commonest method of mining deep ore deposits in Canada. These methods account for almost all the pro¬duction from Falconbridge Mines at this time. The ad¬vantages claimed are: safety-a minimum of open ground; very little supplementary ground support (except rockbolted backs and walls); greater selectivity and flexibility in extraction (horizontal extension of the vein, high grade ore stringers going out into the walls); and high productivity and increased mechanization in larger ore widths. There is some difference of opinion as to whether hydraulic classified tailing back fill provides any great resistance to initial closure of the stope walls. The evi¬dence available suggests that it does not. Addition of portland cement to the back fill just prior to placement has produced a material having the properties of a weak concrete. It is confidently expected that this material will assist greatly in the efficient mining of pillars and generally in giving increased ground support to oppose closure. PRINCIPLES OF STOPE PLANNING BY LONGITUDINAL FLAT-BACK CUT-AND-FILL Again quoting from Spalding: "The practice of con¬trolling output by stopping and starting stopes is bad¬stopes once started should proceed without interruption, and control must therefore be obtained by altering the rate of stoping. It is therefore necessary that faces should normally advance at a medium rate which can be boosted, if required, to give temporarily a greater output. Modern high productivity stoping methods, however, generally aim to achieve and maintain the maximum output possible from a stope. Control of tonnage and grade, therefore, becomes a more complex problem requiring very careful scheduling and organiza¬tion. With the basic method of flat-back overhand stoping using hydraulic fill, it is not usually possible to mine in such a way as to avoid creating pillars." Consider an ore body as illustrated in Fig. 1. It is 274.3 m (900 ft) high by 228.6 m (750 ft) long on strike and averages 4.6 m (15 ft) in ore width. It rep¬resents approximately 907 000 t (1,000,000 st). The ore body could be mined by advancing one long single face through all the levels from bottom to top. A total face length of 228.6 m (750 ft) would accommo¬date five 45.7-m (150-ft) stopes. See Fig. la. Each of these stopes would produce approximately 1814 t/m (2000 stpm) and the ore body would require a period of ten years for complete extraction. A possible im¬provement to this scheme might be to stagger the indi¬vidual stope faces. See Fig. 1b. Problems of hanging wall failure can be reduced by this arrangement. From the point of view of reducing travelway maintenance, it is good practice to employ a stoping system of retreat. Stopes farthest from the shaft are mined first so that on completion, haulage drifts below can be aban

Jan 1, 1982

By William E. Foreman

Editor&apos;s Note: For purposes of uniformity throughout this book, Coal Preparation, 5th Edition, the editors have advocated the use of the U.S.A. Sieve Series of testing sieves (see Table 6-2) as the testing sieves of modem choice. The adoption and use of the U.S.A. Sieve Series and its forerunner, the Tyler Standard Screen Scale, has for over seventy-five years introduced uniformity into reports; has simplified the comparison of results; and has provided control for operation of modem coal preparation circuits. While a full explanation of the differences, and more importantly the similarities of these two measuring systems cannot be stated here, it is felt that a brief clarification of terms is in order. The difference between the U.S.A. Sieve Series and the Tyler Standard Screen Scale Sieve Series is the identification method. Tyler Screen Scale sieves are identified by mesh designation with the width of opening varying by the square root of two, or 1.414 times the width of opening in the previous sieve. The U.S.A. Sieves are identified by openings in millimeters or microns and also arbitrary numbers which do not necessarily mean the mesh count; this series uses width of opening variations in the order of approximately the ratio of the fourth root of two, or 1.189. Tyler Standard Screen Scale and U.S.A. Sieves can be used interchangeably, and to avoid confusion the nameplate of all Tyler Screen Scale sieves show the U.S.A. Number, while all U.S.A. Sieves show Tyler Screen Scale mesh equivalent numbers on the nameplate. INTRODUCTION Sizing is the process of separating mixed mineral particles into groups of particles all of the same size, or into groups in which all particles range between definite maximum and minimum sizes. This chapter applies the modem mechanical screen and the liquid-solid cyclone (hydrocyclone) to the process of sizing solid particles as applied to coal preparation. Screening is a mechanical process that separates particles by size. Modem coal preparation is made possible by efficient sizing. Sizing of coal is defined in terms of the opening through which an individual particle will or will not pass, or as two openings, the larger of which passes the particle and the smaller of which retains the particle. Screening accomplishes sizing by passing coal of a wide range of sizes through a series of screens, each of a decreasing size. The individual screen discharges are then directed into different screen products for subsequent sale or processing. Screening in modem preparation plants is done for the following reasons: 1. Removal of fine coal before a size reduction step. 2. Size separation of the coal into a coarse and fine fraction for sized products or for processing in coarse and fine coal circuits. 3. Recovery of heavy media from heavy media coal cleaning circuits.

Jan 1, 1991

By G. Gregory Bryan

G. GREGORY BRYAN 33.1-DEFINITIONS 1. Engineering-Engineering may be defined simply as the science and art of making materials and energy serve the purposes of mankind. There are many fields of engineering, each with specialized branches. The basic fields are: civil, mechanical, electrical and chemical. Others may be specialized phases of one of these or a combination of two or more of them. Mining engineering, for example, requires a working knowledge of construction, mechanics, electricity, metallurgy, geology, accounting and economics, in addition to mining techniques or methods. Mining geology and geological engineering involve rocks and their structure and the forces that act upon them, and chemistry, physics and mechanics also are important to practitioners in these fields. Metallurgical engineering requires chemistry and physics but also a working knowledge of mechanics and electricity. The fields of engineering with which this section are concerned are those primarily involved in the finding, mining, processing and utilization of the natural mineral resources of the earth. They include geology, mining engineering and metallurgy, each with several specialized subdivisions. 2. The Engineer-The engineer is an individual who has had specialized training to qualify him to practice the profession of engineering. The term engineer is used here in its broadest sense to include such professions as certain applied branches of geology, physics and chemistry, whose practitioners may not have acquired engineering degrees but who, nevertheless, by training and experience in some area of the minerals industry, qualify under the above definition. To become a successful engineer, a person should have certain pronounced aptitudes and preferences, even before embarking upon the requisite course of training. Among the chief qualifications that the engineer-to-he must possess are: (a) a natural inclination toward mathematics and the physical sciences, and (b) intellectual and moral honesty. Among the qualities the student engineer must develop are: (a) independence of thought, open-mindedness, sound judgment and the ability to observe, deduce, plan and apply solutions, (b) the ability to articulate and convince, and (c) the ability to organize and direct the activity of others. These qualities are normally developed through sound education, followed by pro¬gressive experience. 3. The Consulting Engineer-The consulting engineer is an engineer primarily in the business of furnishing specialized advice and engineering services on an intermit¬tent, part-time or occasional basis to clients who do not have continuous or full-time need for those services. He must be able to make and take full responsibility for decisions and recom¬mendations where original thought is required. He should not be confused with the engineer, surveyor or technician who also may furnish services on an intermittent basis, such as regular periodic surveys, the overseeing of drilling programs and other such routine work as may require little original thought and planning. 33.2-THE PROFESSION OF CONSULTING ENGINEERING 33.2.1- QUALITIES REQUIRED A successful consulting engineer must first have, as a minimum, the qualities and preparation of an engineer. In addition, he must have sufficient knowledge,

Jan 1, 1973

By Y. H. Alghamdi, A. D. S. Gillies

"INTRODUCTION Background Diesel Particulate Matter (DPM) has been a critical issue during the last decade in underground mines. This problem is threatening the health of mine workers, as well as the mine’s productivity. It has been confirmed that the diesel engine exhaust is the main source for the diesel particulate matter (DPM), and that it has the highest impact in raising carbon monoxide, carbon dioxide, oxide of nitrogen, and hydrocarbons in the underground environment. Miners who are exposed to this kind of emission are at high risk for developing lung cancer and other lung diseases. The Mine Safety and Health Administration (MSHA) has issued standard rules to regulate diesel emissions in underground mines. The final ruling confirmed in 2008 that the total carbon should not exceed 160 µg/m3. (MISHA, 2008) Monitoring DPM in underground mines has become important to ensure compliance with the MSHA regulations. There are many different types of DPM monitoring systems that are currently being used in underground mines. Two methods are typically used for DPM sampling: the personal sampling method or the direct exhaust sampling method, which takes a direct sample from the engine’s tailpipe. Many mines do not recommend using the direct exhaust sampling method since it involves complex measuring equipment and procedures, as well as it is not regulated in the United States or Canada. Thus, the personal sampling method is the preferred method in many mines when taking DPM samples. Even this method includes many variations, such as the Respirable Combustible Dust (RCD) method and the method devolved by the National Institute for Occupational Health and Safety NIOSH 5040. The NIOSH 5040 method is the preferred method overall because it gives more accurate measurements compared to the other methods. This method does not allow organic carbon sources other than those coming from diesel engine exhausts to interfere with the measurements, which otherwise would lead to overestimating the actual amount of the DPM in the mine environment. An impactor feature in the cassette, which holds the samples in the instrument, characterizes this method. Since the diesel particles are less than 0.9 micron, this impactor does not allow particles more than 0.9 microns to be collected in the cassette filter. The cassette is then analyzed in a lab and the concentrations of organic carbon (OC), elemental carbon (EC), and total carbon (TC) are determined. Lately, the National Institute for Occupational Safety and Health (NIOSH) has developed a new instrument that can use the same concept of NIOSH 5040, but with instant measurements of elemental carbon. It utilizes laser radiation with the cassette so it can give a real time measurement. The real time measurements give instant information about the underground environment that will allow the mining engineer to take immediate action and enact plans instead of waiting for results from the laboratories where NIOSH 5040 method may take several weeks to get the measurements."

Jan 1, 2016

By Meng Wang

"In this study, multiple methods that include geological surveying, laboratory studies, and numerical modeling have been used to study the deformation characteristics of a “deep inclined rock roadway with weak planes” (DIRRWP). Failure of the DIRRWP was found to be triggered initially by the shearing of the rock near the weak plane. The movement of the failed rock resulted in reduced confinement of the deeper rocks, and causing the failure of the supporting units and the severe deformation of the surrounding rocks. Hence, it is concluded that DIRRWP stability is affected predominantly by rocks near the weak plane. Corresponding control strategies for DIRRWP deformation were proposed: (1) greater support intensity is necessary to maintain the stability of the whole DIRRWP; (2) additional support is needed for rocks near the weak plane to reduce the potential of failure; (3) to increase the bearing capacity of the floor strata, as the floor needs to be supported; and (4) a supplementary support technique should be considered for cases of large deformation. The case study presented here indicates that the proposed control strategies can significantly reduce the potential of failure in rocks near the weak plane, thus preventing severe deformation of the DIRRWP.INTRODUCTIONAs shallow seams are mined out, the cover depths in most Chinese coalmines have become increasingly greater. Deep cover in a coalmine is considered to occur when the cover depth exceeds 800 m (Sun, Chen, and Miao, 2018). The loading and failure behavior of rocks in a deep cover setting are quite different from those in a shallow cover setting owing to the increase in in-situ stress, temperature, and hyperosmosis (Lobanova, 2008; Zhang et al., 2013). For instance, experience shows that bulking and dilating deformation in deep roadways is much greater than that which occurs in shallow roadways (Hou and Yang, 2018; Zhang et al., 2016). In addition, anisotropic behavior and thickness deviation of coal measure rocks, combined with the appearance of weak planes (e.g., faults, joints, and coal cleats) in coal and/or rocks, can significantly affect the stability of an underground roadway (Ghazvinian et al., 2013; Sun, Chen, and Miao, 2018). The presence of weak planes can significantly complicate displacement and stress transmission in nearby rocks. Additionally, tensile, compressive, and/or shearing failure in rocks near weak planes can be easily caused by high in-situ stresses in deep roadways; such failures may lead to further severe deformation of the surrounding rocks, resulting in roof falls and rib spalling (Kamata and Mashimo 2001; Wu, Ohnishi, and Nishiyama, 2004). A typical failure of the so-called “deep inclined rock roadway with weak planes” (DIRRWP) is shown in Figure 1. Severe deformation can be found near the weak plane. In this study, a plane or thin layer of rock/coal parting that can significantly reduce the strength of the surrounding rocks—and thus affect roadway stability—is defined as the weak plane."

Jan 1, 2019

By Gerard P. Hand

The use of graphite in industrial applications is one of the oldest in the world. Yet, it is one of the fastest-changing technologies today. Graphite was first used centuries ago in developing writing instruments and as a raw material in crucibles used in melting metal. Today, graphite is used in some of our most demanding industries. These include aerospace, electronics and automotive applications. Graphite usage has historically been split between the natural and synthetic forms. Figure 1 shows the breakdown of natural graphites and synthetic graphites used. The usage of natural graphites and synthetic graphites is split fairly evenly, although the applications are quite different. Figure 2 shows the breakdown of applications using natural graphite. The largest application is the refractory industry. This includes refractory brick and plastics, crucibles, shrouds and other refractory parts. This main application uses roughly 50% of all natural graphite. The refractory industry is an important application for natural graphite. However, the refractory markets have been struggling for several years. As demonstrated in Fig. 3 (Semler, 1996), the production of refractories in tons per year has declined by more than 32% during the last 25 years. Refractory prices have actually risen on a dollar per ton basis during that same period as the product mix has shifted to the value-added side. Raw material and production cost increases, inflation and other related costs have kept profitability at low levels for the industry. Furnace linings that once lasted 1,000-2,000 heats now last 8,000-12,000 heats. This means that tonnage requirements have gone down and quality requirements have gone up for graphite suppliers. Therefore, new opportunities and applications must be developed if the graphite industry is to grow. Synthetics are the other major forms of graphite. There are two forms of synthetic material currently used. One is the traditional type. It is made from the turnings of graphite electrode production. The second was developed in the mid-1970s in response to a shortage of conventional synthetic graphite. This occurred by a breakthrough in thermal technology. This new technology purifies, desulfurizes and graphitizes petroleum coke on a continuous basis at temperatures of more than 2,500° C (4,532° F). This product is called desulfurized petroleum coke (Desulco). It now accounts for about half of the synthetic type material used in the United States. Figure 4 shows the breakdown in the use of synthetic types of material by application. The largest requirement is for use as a carbon addition in iron and steel. This is followed by alkaline batteries and friction materials. Current technological opportunities New opportunities for graphite products are now being developed. These new areas are made possible due to the advances in graphite thermal technology. The ability to purify and modify graphite and carbon products is the key to future growth in the graphite industry. New, higher-purity, more consistent graphite products are leading to enhanced performance in current applications. They are also opening the door to new applications.

Jan 1, 1997

By Donald Whittaker

Severe climatic conditions in mines are normally associated with working at depth. More precisely, it is a matter of working in regions of high strata temperature, since, for example, some coal mines in the Ruhr have strata temperatures at 1000 m which are matched only at 3 times that depth in the South African gold mines. In Britain we are perhaps fortunate in encountering strata temperatures in our deeper coal mines some 10-15 K lower than at the same depth in the hot mines of the Ruhr. Nevertheless the problems of high temperatures and humidities in UK mining operations, which in the past have been confined to a few specific sites, have in recent years become more widespread, and there is every reason to suppose that the trend will continue. Clearly strata temperature is not the only .causative factor. Long ventilation circuits serving districts remote from the shaft must also play a part, while the heavy use of machinery and the rapid, rates of excavation of modern mining methods are obviously important. It is necessary to understand the mechanisms of all such factors, and to measure their magnitudes and fluctuations, if remedial measures are to be properly designed or predictions made which are accurate enough to be of use in ventilation planning. This paper presents the results of a long-term (and continuing) investigation carried out in several mines to measure the major sources of heat and moisture on longwall districts and trunk conveyor roadways. The variation with time of these sources is reported, and the resultant effect on the climate is considered. In their investigations of the heat exchanges occurring in a whole mine Oakes and Hinsley (1955-56)1 listed and estimated the magnitudes of a large number of heat sources. Some of these, namely autocompression of the air in the shaft, diesel engines and compressed air mains, do not apply to the present investigations, which refer to longwall districts with conveyor transport and no compressed air supply. Others, such as oxidation of coal and, desorption of methane, have not been taken into account in the measurements, but are discussed later. One other source, the importance of which was pointed out by Gracie and Matthews (1975)2, is the coal on the conveyor, a form of strata heat emission, but one which is strongly dependent on the rate of coal production. The heat emissions measured in the present investigations were the energy consumed by electrical machinery, the heat conducted from the strata and the heat lost by the conveyed coal. Most of the emitted heat was removed from the district by the ventilation air, and was measured by continuously recording the enthalpy of the air at fixed stations. The heat removed by service water, supplied for machine cooling and dust suppression, was also measured. Account was of course taken of the relatively minor changes in level around the districts, in terms of the potential energy gained or lost by the coal and by the ventilation air (autocompression).

Jan 1, 1980

The latest resource estimate for the Caspiche gold/silver/copper project in the Maricunga region of Chile was released by gold explorer Exeter Resource Corp. The resource is low grade, but bulk minable and drilling to date has shown a NI 43-101 compliant inferred resource of 1,117 Mt (1.23 billion st) grading 0.55 g/t (0.016 oz/st) gold, 3.81 g/t (0.111 oz/st) silver and 0.22 percent copper. Exeter said these findings make the in situ metals amounts of 609 t (19.6 million oz) of gold, 4.4 kt (137 million oz) of silver and 4.84 billion lbs of copper, which the company relates as 33.7 million gold equivalent ounces using assumed metal prices of $23.33/g ($800/oz) for gold, 35 cents/g ($12/oz) for silver and $2/lb for copper. The deposit is almost midway between Barrick?s Cerro Casale gold deposit and Kinross? Refugio Mine. In a statement Exeter chairman Yale Simpson said ?This updated resource estimate places Caspiche firmly among the world?s largest gold-copper discoveries in recent years and clearly supports our decision to aggressively drill the property, despite the global economic decline. ?We expect that the very substantial copper and silver credits will be extremely important, and will markedly impact the potential viability of Caspiche. Specifically, using metal prices of US$2/lb copper, $23.33/g ($800/oz) gold and 35 cents/g ($12/oz) silver, the 4.84 billion lbs of copper and 4.2 kt (137 million oz) of silver are equivalent to 376 kt and 62 t (12.1 and 2 million oz) of gold, respectively. ?Our recent drilling has defined what appears to be a coherent higher grade central zone associated with a definable early stage diorite intrusion. We expect this zone will be important in developing various mining and economic models for the project,? Simpson said. Exeter is still drilling the deposit to expand the re-source estimate further to the southeast and at depth and to upgrade the higher grade central zone to the indicated category. Simpson said, ?Exeter intends to enhance the value of Caspiche by continuing to de-risk the project through ongoing exploration, metallurgy, engineering, water and environmental studies. These studies will continue through the next 12 months and will lead to a conceptual development study. A timeline for the release of that study is being formulated. The project budget for the next 12 months is C$14 million.? Mineweb reported that Caspiche is likely too big of a deposit and at too high of altitude for Exeter to handle alone. Exeter will be looking to other majors, or perhaps the Chinese, to help develop this huge orebody. But, in the meantime, it also has its high grade Cerro Moro deposit in Argentina, which could be brought to production relatively quickly and at relatively low cost.

Jan 1, 2009

By Richard Dunmore

Most coal seams in the United Kingdom contain firedamp, some of which is released into the ventilating airways as the strata around a working relax. The close sequence of seams in carboniferous strata is usually such that seams in adjacent roof and floor strata are also disturbed and potentially large quantities of firedamp can enter the workings from these sources. Since emission depends on the disturbance rate of the seams, higher coal production rates associated with increased mechanisation in the colliery accentuate the problem. Firedamp concentrations in UK underground workings are subject to statutory regulations1. To enable production to continue uninterrupted and in safety the problem is tackled by ventilation engineers by the provision of adequate ventilation in the airway and by the technique of firedamp drainage. In the UK the latter usually takes the form of cross-measure boreholes in roof strata, and increasingly also in floor strata, drilled from the gateroads, especially from return airways. In this way firedamp emitted from disturbed seams in adjacent strata can be extracted directly to the surface or to some safe position outbye, and the airways of the working district are not contaminated. Despite the use of firedamp drainage, considerable quantities of firedamp enter the ventilation air- stream and the concentration levels may be seen to be related to the coal production activites in the district. Figure 1 shows a typical recording of methane concentration obtained by a BM1 methanometer at a position outbye in a return gate where firedamp drainage was operating. The general rise in level throughout the working week may be seen, together with peaks occurring during production shifts and the gradual decay of emission over the weekend. In view of the fact that ventilation and firedamp drainage contribute a considerable proportion to the operating costs of the colliery, it is essential that the ventilation and firedamp drainage requirements be calculated as accurately as possible at the planning stage of the working. In the past the ventilation engineer has had to rely heavily on experience in neighbouring workings but with the development of mines in previously unworked areas there is clearly a need for a methodical approach to meet the requirements of planning. The work described in this paper represents such an approach and it is hoped that successful development and application of the method of prediction will lead to the widespread adoption of a routine procedure. The current development of a method of prediction of firedamp emission into longwall workings is based on the theoretical treatment of gas emission from coal seams proposed by Airey2 in 1971, which was based on his study of gas emission from broken coal 3 which was published in 1968. In common with other experimental methods of prediction, the method enables a calculation to be made of the total make of gas emitted from all sources of gas in the strata per tonne of mined coal.

Jan 1, 1980