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By Atsushi Sainoki

Induced seismicity is deeply related to various engineering projects entailing anthropogenic in-situ stress change in deep underground. Hence, there is a persistent demand for estimating and mitigating the risk. Significant efforts have been made in simulating induced seismicity with a variety of numerical simulation techniques; however, there are still discrepancies between characteristics of induced seismicity numerically simulated and observed in the field. One of such discrepancies is the simulation of triggered seismicity defined as seismic events caused by an infinitesimal amount of anthropogenic stress change, which is frequently observed in geological structures. The present study aims to develop a methodology to simulate triggered seismicity whilst reproducing a complex stress state in a fault zone. For the purpose, a discrete fracture network within a fault damage zone is first generated. Then, an equivalent continuum model, in which each zone has a different anisotropic compliance tensor, is constructed by employing the crack tensor theory. A fault core is simulated at the center of the model with interface elements. Using the equivalent continuum model, an initial stress condition is simulated with the boundary traction method, which results in a very complex stress state with stress anomalies due to the stiffness anisotropy and heterogeneity. Subsequently, the effective normal stress on the fault plane is decreased in a stepwise manner to induce slip. The result demonstrated that many seismic events are induced on the fault plane with a slight decrease in the confining stress, thus being regarded as triggered seismic event. Further analysis on this seismic activity shows that the magnitude of triggered seismic events is inconsistent with the amount of shear stress drop, implying that a relatively small triggered seismic event could produce a large stress drop, generating locally intense ground motions and thus inducing severe damage to the nearby rock mass.

Apr 26, 2022

By J. R. MacDonald

MINING literature contains very few descriptions of operations producing as little as 25 to 30 tons of ore per day. The purpose of this paper is to discuss the gradual development of a mining method for such a small-scale operation at the No. 2 mine of Central Patricia Gold Mines, Limited, and in doing so to review several systems of stoping applicable to narrow ve1ns. MAIN ECONOMIC FACTORS TO BE CONSIDERED (1) Overhead.-The outstanding feature of most mines operating on a single narrow vein is chat of small daily tonnage. The reason for this low output can be readily understood. The strictly limited ore-reserves present a condition unfavourable for a policy of large initial expenditures, with the result that the capacity of the plant provided must be governed by the? amount of ore in sight at the time when mill construction is planned. This is immediately reflected in the cost per ton, there being relatively few tons over which to distributed the more or less static charges, such as hoisting, heating, supervision, administration, pre-production expenses, and other write-offs. (2) Development.-In mining this type of orebody, the main factor contributing to high costs is unquestionably the small tonnage of ore made available by a given amount of development. Thus, for example, from the data tabulated below it can be seen chat, assuming a 12-in. vein and a drifting cost of $9.00 per foot, the development charge for this item alone would be 71 cents per ton. Add to this the charges for raising, cross-cutting, shaft-sinking, and pumping, besides charges arising from development in waste, and the figure will be very considerably higher than 71 cents.

Jan 1, 1941

By Jean-Pierre Birat, Ana-Maria Iosif, Denis Ablitzer, Olivier Mirgaux

"Integrating environmental considerations into the product traditional process design is now the major challenge for steel industry. Life Cycle Assessment (LCA) is nowadays considered as an appropriate method for assessing environmental impact and selecting new technologies to reduce CO2 emissions for steel industry.In this paper we propose a new methodological concept which combines LCA thinking with process simulation software in order to carry out the life cycle inventory of classical steelmaking process.Using Aspen PlusTM software, a physicochemical model has been developed for the integrated steelmaking route. This model gives the possibility to carry out life cycle inventories for different operational practices in order to optimise the use of energy, to calculate CO2 and other emissions and to control the mass and the heat balances of processes.It is also shown that such approach can be used to design and assess new technologies for steelmaking without large industrial application.IntroductionDuring recent years, it was recognised that CO2 emissions are the major factor of the global warming effect. In order to meet Kyoto requirements, the steel industry is embracing the challenge of sustainable development by improving its competitiveness and economic success while reducing its environmental impacts. Integrating environmental considerations into the traditional product design process, for powerful eco-efficiency, is now one of the major priorities for steelmakers.The European Steel Industry has measured up to this challenge by creating ULCOS (Ultra Low CO2 Steelmaking) consortium that has taken up the mission to develop breakthrough steelmaking process with the potential to reduce CO2 emissions (Eurofer, 2005). In order to develop technologies to reduce emissions, it is necessary to assess the environmental impact of the classical steelmaking route (coke plant, sinter plant, blast furnace, basic oxygen furnace, continuous casting and hot rolling)."

Jan 1, 2009

By John Nagy

The effect of low percentages of methane (0 to 5 pet.) in an air atmosphere on spark initiated explosions of coal dust was investigated. With low dust concentrations (0.1 to 0.4 oz./cu, ft.), small amounts of methane increase the explosion pressure; with larger dust concentrations (0.5 to 2.0 oz./cu. ft.), small amounts of methane decrease the explosion pressure. The rate of pressure rise of a coal dust explosion increases with increase in methane concentration; about 3 to 4 pct. further addition of methane yields a decrease in the rate of pressure rise. Convection currents created by the air or gas blast used to disperse dust produce an increase in explosion pressure and rate of pressure rise. The minimum explosive concentration of the coal dust decreases about linearly as the concentration of methane increases from 0 to 5 pct. The explosion pressure increases linearly as the initial pressure within the chamber increases. Similarities between coal dust and methane explosions are indicated, INTRODUCTION During coal mining, methane occluded in coal or adjoining strata is liberated into the mine atmosphere,4 Normally, the methane is diluted and swept away by the ventilating air. Under some conditions, through accidental disruption of the ventilation system, sudden outburst, or because insufficient fresh air is supplied to a working place, low percentages of methane may be present during mining operations. If it is permitted to accumulate, methane itself presents an explosion hazard; its limits of explosibility arc approximately 5 and 15 pct. The effect of low percentages of methane on the explosibility of coal dust was studied to provide a broader understanding of the hazards of ignition and explosion and thus to facilitate development of control

Jan 1, 1961

By P. Lenegan, R. Masethe, R. I. L. Ferreira

Underground mining in high rock-stress environments can induce unanticipated dynamic rockmass deformation, posing a serious risk to the workforce. Data collected by seismic networks are imperative for the design-as-you-mine process. Often an inappropriate assessment of the seismic hazard is done when unnecessary data are included in analytical techniques, particularly short-term time frames. Depending on the methodology or application of data analysis and interpretation, some of the information may not be relevant and even detract from the objective. By using the spatial event clusters associated with active workplaces to define the shape and size of the polygons for data selection for subsequent analysis, a problematic subjective component can be eliminated. The mining of tabular reefs in South Africa, be they gold- or platinum group metals, is very extensive, and underground workplaces are often widespread. Production stoppages are very costly, especially for marginal mines, so there is just cause not to evacuate an entire operation’s workforce following the occurrence of a large mining-induced seismic event. A rationale based on the damaging peak particle velocities (PPV) of the past seismicity can exclude the unaffected underground personnel and allow a quick orderly withdrawal where required. Micro-deformation may precede rockbursting, which can occur on-shift close to the workface. A mine-wide array of geophones suitable for the accurate location and analysis of seismicity may not be able to detect these higher-frequency emissions. Inexpensive but fit-for-purpose accelerometers, providing real-time, continuous measurement of acoustic rockmass deformation near the workplace, integrated through underground and surface communications networks, are shown to record ‘rock-talk’, clearly distinguishable from the instrument background noise, which can be in the order of hours leading up to an imminent damaging seismic event.

Jan 12, 2025

By Jr. Watts, B. K. Cantrell, K. L. Bickel, J. F. McDonald

"A field trial was conducted at an underground gold mine to evaluate diesel fuel derived from soybean oil (soy-methyl-esters or “biodiesel”) as an alternative to commercial, low-sulphur, number 2 diesel fuel (D2). A 2.7 m3, 18-ton load-haul-dump (LHD) unit, powered by a naturally aspirated, indirect injection (pre-chamber), 100 kW diesel engine was used as the test vehicle. The LHD was equipped with a diesel oxidation catalyst. The LHD operated satisfactorily on the biodiesel fuel. Operators observed a slight decrease in power and an absence of black smoke.Diesel particulate matter (DPM) concentrations were monitored for a six-week period while the LHD was operated with either the biodiesel fuel or the D2 fuel. DPM concentrations were measured in the mining section’s clean air intake, exhaust and on the LHD. Reduction in DPM emissions was determined from the measured difference in DPM concentrations between the clean air intake and the exhaust. The biodiesel fuel reduced energyspecific DPM emissions approximately 75%, from 3.6 ± 1.8 g/kW-hr to 0.86 ± 0.27 g/kWhr. Time-weighted average DPM exposure, measured near the LHD operator, was reduced from 0.67 ± 0.13 mg/sm3 to 0.28 ± 0.12 mg/sm3, or approximately 58%, when the biodiesel fuel was used. IntroductionDiesel exhaust consists of noxious gases and diesel particulate matter (DPM). Recent attention has focussed on the potential carcinogenicity of DPM and the potential health impact on miners. The National Institute for Occupational Safety and Health (NIOSH) recommended that whole diesel exhaust be regarded as a “potential occupational carcinogen”, and that reductions in workplace exposure would reduce cancer risks (NIOSH, 1988). The International Agency for Research on Cancer (IARC) declared that “diesel engine exhaust is probably carcinogenic to humans” (IARC, 1989)."

Jan 1, 1997

By B. Hines

Grinding efficiency in a vertical stirred mill is affected by a wide range of process and mill configuration variables. Of these, the media fill, size, and intrinsic density have a significant effect on power draw and grinding performance. Media-related factors that influence grinding performance include, e.g.,a higher media filling providing additional particle breakage sites, or a lower bead density decreasing the running torque at a given media filling. It has been estimated that the grinding media can cause up to a 30% change in the specific energy required to grind. FLSmidth has generally recommended that the intrinsic density (specific gravity) of the grinding media be more than triple the slurry density. This density difference helps to ensure that sufficient stress intensity is generated in the mill to provide particle breakage. It also reduces the chance of media ?float? if there is a spike in either slurry density or slurry flow rate. The VXPmills have historically operated at slurry density values between 1.2 and 1.5 kg/L, which translates to nominal bead intrinsic density range between 3.6 and 4.5 kg/L. The current trend towards regrinding bulk flotation concentrates to even finer sizes, to improve recoveries of fine grained minerals, has provided motivation to consider closed-circuit operation of stirred media mills. The slurry density in this application is typically higher than in open-circuit operation due to the higher percent solids in cyclone underflows. An additional effect is that closed-circuit operation increases the potential for media float due to higher flow rates as slurry is recycled to the mill. The normal progression in extending the operating range of existing equipment is to carry out laboratory testing before pilot-scale and plant-scale trials. The results presented in this paper examine the media density to slurry density ratio and its effect on grinding performance.

Feb 23, 2014

By J de Witte

When used in suitable ground conditions, rock bolts can have advantages over other permanent tunnel lining solutions. Permanent rock bolts offer flexibility in tunnelling profile and rapid rock bolt installation can enable time and cost savings. There are also disadvantages when considering the suitability of rock bolts for permanent support installations. These generally concern achieving the specified durability requirement over the specified design life. By investigating alternate approaches to rock bolt installation, quality assurance (QA), monitoring and design life it may be possible to alleviate some of these concerns.The design and construction benefits that can be realised through the use of rock bolts are primarily related to reducing project program time. This time reduction could be realised if rock bolts are considered as a permanent solution, with the installation being completed as a part of the primary excavation cycle. Any reduction of the program could allow direct and indirect cost savings and in the case of a major project, allow the revenue stream to be realised earlier.Currently ambiguity exists where the treatment of rock bolts as permanent support components is inconsistent with that of (often identical) ground anchors, as utilised in stabilising cut and cover structures. The ground support design process, implementation and testing are generally similar, though the acceptance of rock bolts as long-term support elements is treated differently to ground anchors in technical specifications.This paper discusses some of the benefits available through the use of rock bolts. It offers different solutions to the concerns surrounding durability, including alternate approaches to installation, Quality Assurance practices and monitoring solutions. The selection of an alternate design life is also investigated as another means of alleviating concerns around long-term durability. The paper will also provide an indication of savings that could be expected through the use of rock bolts as part of a long-term support system.

Mar 8, 2011

By M Gwynne

McArthur River Mine is located approximately 50 km south-east of the NT township of Borroloola and approximately 120 km from the Gulf of Carpentaria. The mine hosts a series of layered shallow dipping tabular orebodies, with current reserves of approximately 22 Mt, grading 13.6 per cent zinc, 6.3 per cent lead and 65 g/t silver. Owing to economic considerations, the lower most orebody (No 2 orebody) has been selected for initial mine production. Following the completion of the No 2 orebody, mining will transfer to the lower grade and metallurgically more complex No 4 orebody. The mine employs a total of 235 people throughout the underground and surface areas. The majority of the workforce work a fly-in/fly-out, seven-day- on/seven- day- off roster with 12-hour shifts. Presently, a further 18 contractors are involved in trial mining areas. An apparent dip room and pillar technique has been selected as the primary mining method. Officially opened in September 1995, the current production rate of 1.2 Mt per annum will eventually escalate to 1.5 Mt per annum. Though most production is achieved through the use of conventional loaders, trucks and jumbos, the recent development of methods utilising scrapers, hand held machines and modified jumbos has produced encouraging results in steeper dipping areas of the mine. Initial production has been hindered by several factors including extensive and largely unpredictable faulting, worse than expected ground conditions and difficulties with some mining equipment. Compounded by the relatively thin 3 m ore seam, initial mining resulted in high dilution and poor production. This paper describes the development of unique planning and production systems that were evolved in order to overcome these problems and achieve the targeted production rates and head grade.

Jan 1, 1998

By R van Wyk

In today’s mining environment the use of heavy mechanical devices and other high-tech electronic devices like two-way radios, cellular phones and even automated mining systems, form a vital part of the daily operations of a mining site. These human-made devices and some of the naturally occurring events, like lightning and gamma radiation, can have negative effects on electronic detonators and the associated control equipment.On larger mining operations the necessity to move the point of safety for blasting further away from the blast area, introduced remote firing via radio frequencies for electronic detonator systems. Other influences on electronic detonators can be abuse or ignorance. To ensure the safety of the end-user, an electronic detonator must be resistant to some common mistakes that users can make. One of main concerns here is resistance to static discharge or incorrect wiring. Despite the daily use of electronic detonator solutions, scepticism and concern still exists amongst those who believe the system raises the risk of a premature initiation of the blast through some form of interference trigger.This paper will help users understand how electronic detonators and the associated control equipment protects itself from external influences. Focus will be on protection schemes to counter static and over-voltage discharge and methods to ensure protection of secure wireless radio links and detonator communications. With well engineered communication systems and solid static and over-voltage protection, an electronic detonator system can be more reliable and safer than other detonator systems. The following main points will be discussed:what makes a detonator an electronic detonator,static electricity and over-voltage protection,the general influence of electromagnetic energy on the detonator,the influence of lightning on detonators,the protection of radio communication’s links as used by remote firing equipment, andthe safety protection features built into remote firing equipment.

Nov 8, 2011

By A G. Shellsh, R J. Henham, J R. Foley

The Porgera Gold Mine, west of Mt Hagen in PNG is a large open pit and underground operation currently producing gold at the rate of almost one million ounces per year. This deposit has produced some 8.8 million ounces to-date and has current reserves of 103.4 million tonnes @3.2 g/t Au. The Porgera Joint Venture has invested over $A1 billion in infrastructure in this project. Technical management of this operation involves the interrogation of some 700 000 metres of diamond drilling, some 50 000 blastholes per year, and very large volumes of laboratory, geotechnical, environmental and production reporting data. Over the past seven years this data has been consolidated into a single centralised, relational database which now acts as the single reference point for all the technical application systems used for surveying, grade control, production reporting, deposit evaluation and modelling. Data transfer to application programs and management systems is controlled by the centralised database, this process at times involving a two-way, real-time, binary data transfer. The benefits of a specialised database application include security and integrity of the data asset, ease of access, and performance in a multi-user environment. This replaced fragmented, inconsistent data sets stored inefficiently as individual, unindexed, variably formatted files associated with the different application programs. Most importantly, the database has allowed mine management at all levels to make timely, confident investment, cash flow and optimisation decisions as it provides the primary link between the deposit and operational data sets on one hand, and the application programs and management and performance measurement systems on the other. This paper discusses the economic value and importance of technical data in terms of the resource asset and the means by which this data is being managed and used at Porgera.

Jan 1, 2000

By C Zhang, I Canbulat, G Fan, D Zhang, C Wang

A clear understanding of the intensity of coal and rock burst is the basis for future research on burst mechanism and control. There has been a wide variety in the description of coal and rock burst damage at different countries using multi-methods, which makes the common understanding even more challenging. Therefore, this paper presents a broad review on the quantifications of intensity of coal and rock burst and seismic risk around the world, considering critical factors including seismic energy, magnitude, peak particle velocity and ejection velocity. It is found that the seismic energy of coal and rock burst can widely range from 0.002 MJ to 60000 MJ, with equivalent Richter magnitude from -1 to 4. The daily seismic energies can be used for seismic risk potential assessment, with energy levels from 0.1 KJ to 10 MJ, which is much lower than the burst sources. The magnitude of the seismic events can be also estimated based on the weight of damaged coal mass and human senses of vibration and sound. For the seismic intensity represented by peak particle velocity (ppv), it keeps constant within fault radius and attenuates with a speed of 1/r4 and 1/r in near and far field respectively, thus it is possible for a small seismic event to induce considerable ppv causing severe damage if the epicentral distance is small enough. There has not been a strong correlation between the ejection velocity and ppv, as seismic energy is only one of the energy sources that contribute to the kinetic energy of burst damage. CITATION:Wang, C, Zhang, C, Canbulat, I, Zhang, D and Fan, G, 2018. Intensity quantification of coal and rock burst risk – a review, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 223–233 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

By D. Harrington, M. W. Von Bernewitz

Sections 120 and 121 of Article XI. Fire Control, of a proposed Mine Safety Law in preparation, to be used as desired and when approved by any State, contains these provisions: After the adoption of these regulations, no mine operator shall erect any nonfireproof structure within 50 feet of any mine entrance. or airway. Every operating mine having nonfireproof structures within 50 feet of any mine entrance or airway, except headframes necessary for hoisting, shall have such structures fireproofed or removed and replaced with fireproof structures within six months of the adoption of these regulations. Before reopening, mines idle at the time of the adoption of these regulations shall have all non- fireproof structures fireproofed, removed, or replaced with fire- proof structures as required by the foregoing sentence. Upon receipt of the written statement of a mine operator that the fireproofing of existing structures is impracticable, the chief mining engineer (of the State commission) shall satisfy himself as to the abouracy of the statement. If he concurs that fireproofing is impracticable, and if at the same time he holds the opinion that adequate protection can be provided otherwise by such procedure, he shall have authority to issue a written permit for the substitution of fireproof door at or near portals of mine openings and collars: of shafts. Such doors shall fit as nearly gas-tight as possible. and shall be installed in such a way that the mouths of mine open- ings can be closed readily from outside of structures. The effec- tiveness of such permit may be made contingent upon the operator's continued compliance with any additional precautions which are necessary in the chief mining engineer's opinion to protect the lives of miners.

Mar 1, 1932

"In the conceptual stages of engineering, benchmarking and empirical ratios are commonly the only references for estimating the production rate of greenfield projects. For underground mining projects, the lack of similar cases often makes the comparison difficult due to each project’s unique geology, rock quality or other singular characteristic. This paper discusses an analytical alternative for determining the maximum achievable production rate for underground stoping mining by modeling production as a continuous process. The results show that long-term maximum capacity depends exclusively on the achievable development rate and the development intensity of the design. In operations constrained by backfill, the development rate could be limited by the backfilling rate. Extraction rate and stope size only have impacts on capacity estimates during the ramp-up period.These findings allow shaping of the maximum capacity profile of a project. They can be used as a tool for quick assessment of conceptual capacities, and sensitivities of mining rates. Benchmarking tests compared favorably with projects, mines in operation and scheduling exercises. IntroductionA key consideration in the value proposition of a mining project is its capacity or production rate (Tatman, 2001). At a conceptual level, it is common practice to refer to empirical references (Taylor, 1986; Long, 2009), rules of thumb (de la Vergne, 2014), benchmark ing studies or consultant experience to estimate or validate a target value. There is, however, a lack of analytical methodology to guide an early estimate of the maximum capacity.Araneda and Gaete (2004) presented an interesting mathematical approach to estimate the capacity of panel caving mines. Using a continuous model, they confirmed that the mine capacity is essentially determined by the rate of area incorporation, in terms of area per period, and ore column height. This simple relationship has been confirmed by multiple detailed scheduling works"

May 1, 2016

By C. G. Hunkin

In situ, or in-place leaching is concerned with the extraction of metals from natural mineral disseminations without removal from their geologic location which, in many cases, amounts to the reversal of the deposition process accompanied by a presumed increase in velocity of the processes, if the project is to be profitable. Solution mining processes involve removal of minerals such as sulfur, common salt, and potash, generally leaving cavernous voids, in contrast to in situ leaching which term is applied to the selective dissolution of a mineral from its gangue without sensible disturbance of the formation. Where such a process can be operated, the advantages are manifold and have been so recognized by miners from the earliest days of mining. Recent technological developments have encouraged a reassessment of the status of in situ leaching, and it now appears possible to extend the application of the method from a few isolated special situations to many uranium and copper deposits. Anaconda's interest in the techniques stems from long experience with vat and heap leaching, 'dump leaching, and natural leach activities associated with several established mines. Increasing activity through-out the industry has been noted, particularly in leaching low grade pillars and the peripheral zones of operating mines, often by the introduction of leach solutions via boreholes drilled from surface. Reactivation of dormant mines] using similar techniques has been successful. Over the long term, environmental improvement activity will lead to excess sulfuric acid production, even to a disposal problem. The cost of leach solutions and reagents must be reexamined in this light. The mineral industry now classifies as uneconomic deposits of readily soluole minerals whose potential value, on a per acre basis, may De several times that of many currently profitable oil field reser-

Jan 1, 1971

By B Roache

Ground support installation into the face of development headings has become an accepted practice at many Australian hard rock mines. Ground support is installed in the development face to reduce personnel exposure to rockfall while conducting work tasks such as marking up, installing lifter tubes, charging, and face sampling. The ground support system installed in the face is typically friction bolts and mesh. This support system can be effective at reducing rockfall occurrence, but it also increases the cost of development per metre, in terms of capital for the ground support, jumbo and operator capital allocation and the installation time slowing each drill and blast cycle. Other issues associated with meshing and bolting the development face include management of rock bolts and mesh sheets in the fired dirt of every cut, damage to services such as vent duct, geometry issues when installing ground support onto irregular development profiles and operator error. Fibrecrete installation onto the development face for support is an option considered by many mines, but it can be difficult to manage due to the removal of visual indications of misfires and slowing the mining cycle due to the wait time for fibrecrete curing. The use of Thin Spray Liner (TSL) products in underground hard rock mines has been relatively infrequent in the past, mainly due to issues with cost, health and safety, and liner strength. Kanmantoo Mine investigated the use of TSLs for the purpose of development face support, while collecting data on the costs and practicalities associated with their use. Several TSL products were trialled and a preferred TSL product was selected for use as face support. Cost savings, drill and blast cycle timing improvements and other benefits were achieved while meeting the requirement to reduce rockfall risk.

Mar 29, 2023

By Timothy J. Matthews, Ted Klemetti, Timothy J. Batchler, Daniel J. McElhinney

Unplanned tailgate collapses and difficult mining conditions in U.S. longwalls have necessitated the use of innovative tailgate support systems. Severe ground movement induced by longwall loading has caused operators to utilize secondary support systems that can match yield and strength requirements with expected ground reactions. Standing support performance requirements depend on the nature of the ground deformation behavior, and therefore require a good understanding of the interaction between the support and ground to achieve an optimum roof support system. Mining operations often employ standing support strategies that are unique to their mine. A support system that functions successfully in one situation could fail in another situation. It is important that the design and installation of the standing support systems provide the highest probability of preventing roof falls. Pumpable roof supports are currently being used to provide a safe working environment for longwall mining. A full understanding of the impact of these performance characteristics is necessary to optimize the support application and to provide a foundation for making further improvements in the support performance. Optimization of roof support applications helps to ensure the protection of mine workers by preventing roof falls due to inadequate support design. The load capacity and stiffness characteristics of a 27-in.-diameter pumpable support system was evaluated using the National Institute of Occupational Safety and Health (NIOSH) Mine Roof Simulator (MRS). Previously, tests to evaluate the performance characteristics were performed on a single pumpable support. The goal of this new research is to evaluate the performance characteristic effects of the pumpable support system and compare it to an individual pumpable test examination. This research of a pumpable support system configuration simulates the current installation practices in U.S. coal mines.

Jul 1, 2023

By N. Razavinia

Most mining and metallurgical industries produce relatively large quantities of low grade waste heat which they vent, for the most part, to the environment. To date, little is being done to recover some of this heat and transform it from low grade to a high grade energy form such as electricity. A few decades back this was also true of paper, plastics and metals. Recycling was not a prominent activity. Today, we all know that it is and that recycling is a must for sustainable development. To this end, we feel that we are now at that point in time when the recycling of waste heat to electricity must be given serious consideration. By recycling waste heat one can achieve at least one if not two of the following benefits: 1) reduction in the carbon footprint ? by improving the overall energy efficiency of a process, and 2) production of electricity at an economically viable cost. We have developed a process that captures, concentrates and converts (3 C?s) waste heat to electricity. The process comprises 3 distinct interacting units. The first unit captures and concentrates waste heat using heat pipe technology. The second unit converts some of the concentrated heat to mechanical work which can then be readily converted to electricity. This unit is based on the classical piston/cylinder engine arrangement that is widely used by consumers to power systems with compressed air, for example. The third unit of the process dissipates the heat that the engine cannot convert to high grade energy. It also uses heat pipe technology to achieve this. The combination of the 3 units forms a unique system for recycling waste heat. This paper describes the system and it presents some results that show how the system can be made to be economically viable.

Aug 1, 2013

By James J. Dowd

The investigation to evaluate the reserves of coking coal in the United States suitable for the manufacture of metallurgical coke is being made by the Bureau of Mines in three parts: (1) To estimate known (measured plus indicated) recoverable reserves of all coking coal; (2) to upgrade marginal coals through effective preparation; and (3) to study the carbonizing properties of coals and coal blends not now widely used for metallurgical coke making. This is the fifth of a series of reports giving the results of studies by countless under part (1) of the investigation. (See appendix.) This report covers Westmoreland County, Pa., which comprises parts of New Kensington, Freeport, Elders Ridge, Pittsburgh, Greensburg, Latrobe, New Florence, Johnstown, Brownsville, Connellsville, Donegal, and Somerset quadrangles. (See fig. 1.) A short resume" of the chemical and carbonizing properties of the coals are included. Data on the coal beds in this county were obtained from mining companies, land companies, individual landowners, and other sources of authentic information. A base map for each bed in each quadrangle was made to the scale 1 inch equals 1,200 feet. Maps of mines, locations of drill holes, bed and total coal thicknesses, and the outcrop of the bed were plotted on the maps. With all available data plotted, isopach lines were drawn to limit areas of known unmined reserves in beds up to 14 inches thick, 14 to 28 inches thick, 28 to 42 inches thick, and over 42 inches thick. These areas of coal reserves also were divided into "measured" and "indicated" categories. All areas in each thickness range and in each category, mined-out areas, areas excluded from the estimate, and areas outside of the outcrop were measured by planimeter on the base maps. Estimates of total reserves 14 inches and more thick and maps for individual beds were prepared from these data.

Jan 1, 1951

By Emory Smith

The preface of the United States Bureau of Mines Manual of First Aid Instruction contains the following paragraphs: One of the most important inquiries that Congress has authorized the Bureau of Mines to conduct relates to health condi- tions among those engaged in the mineral industries. Investigations made early with a view to bettering these conditions demonstrated the need of prompt care of injured miners, as, under the difficulties inherent to mining and other mineral industries; wounds seemingly unimportant, if not treated promptly, may become infected and possibly cause permanent crippling or even death of the porach. In the past many seriously injured persons have been treated by fellow workers ignorant of approved means of treatment and transportation and have been handled in such a manner as to cause suffering and to accentuate injuries to such a degree as to bring about permanent disablement. Accordingly, the need of general instruction in first aid became apparent. It must be recognized that the conditions in and about the mines, mills, smelters, quarries, oil fields, etc., usually present difficulties not found in large industrial centers where experienced surgeons are available and hospitals close at hand. Many mines are isolated and, when an accident occurs several miles underground, there may be a serious lapse of time before the injured person can be brought to the surface and taken to the hospital or given attention by a sur- geon who may have to come a long distance. Under such conditions it has been frequently demonstrated that, if effective first aid is not given, the injured person may die or become permanently crippled... First-aid instruction has been found to be an excellent aid3in pre- venting accidents. In addition to the immediate relief of injured persons, it has been found that most persons while in the act of practicing dressings for

Apr 1, 1931