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By Raymond Wilson

GRINDING costs are an important item in cement manufacture, and the cost of power is one of the large items in grinding costs. Even where power is of secondary importance, cost items dependent on mill capacity are influenced by the efficiency with which the mill uses power. Tube mills typify the grinding mills most generally used in cement plants, though other types are in use. Tube mills have been widely criticized as being extremely inefficient, and this feeling regarding them has led to many attempts to design more efficient grinding equipment. Other types may be an improvement over tube mills, but the improvement has certainly not been revolutionary, at least as applied to clinker grinding. The very fact that there is doubt regarding the relative efficacy of various types of mills is an indication that an adequate basis for an analysis of performance is lacking. The development of methods of analyzing mill performance has been hampered by the lack of a suitable method for measuring the work actually done in pulverizing. There has been no acceptable standard to which mill performance might be compared. As a consequence, it was not known whether the margin of possible improvement in pulverizing operations was large or small. Even direct comparison of different types of mills was seldom possible because most plants had only one type of mill for each type of service. Comparisons between plants were difficult because of differences in the grinding characteristics of various clinkers. Some recent developments complete a chain of circumstances that place the cement industry in a particularly fortunate position for the study of its clinker-pulverizing operations. Modern investigators of grinding phenomena have generally agreed that the best measure of mill performance is the production of new surface. For most materials this measure of performance is largely theoretical and has a rather remote connection with the practical aims of the grinding operation. In clinker grinding, on the other hand, surface area is intimately connected with the properties conferred on the cement by grinding. Moreover, the determination of surface area is rapidly coming into routine use as at test

Jan 1, 1937

By R. Phillips

The low concentration of diamond in workable deposits requires very large samples for grade estimation, and results, even then, may be erratic and difficult to interpret. Although the Poisson probability distribution cannot accurately describe the occurrence of diamond, it is suggested that it gives a better estimate of grade than conventional methods, even from single samples. Examples illustrate the simple calculation required.

Dec 1, 1971

By Z. Smieszek, Z. Szota, P. Gajowski

"A new method for measuring the total quantity of suspensions contained in the electrolyte during copper electrorefining has been developed by the Institute of Non-Ferrous Metals, Gliwice. Electrolyte purity, and the suspensions content in particular, is of essential importance for the quality of electrolytic copper produced during the process. Suspensions contribute to the increase of slime impurities in electrolytic copper and increase losses in noble metals. Up to now, the laboratory method has been used for the· process control which provides analytical results with considerable time delay. Therefore, it is useless for the process control.The measuring method developed enables continuous measurements of a suspensions content directly within the stream of the flowing electrolyte. The principle of measurement is based on changes in optic properties of the electrolyte with varying suspensions content.1. IntroductionIn copper electrorefining, it is essential that the final product quality be maintained. If suspensions are present in electrolyte, the content of slime impurities in electrolytic copper, (such as Pb, Ag, Bi, Sn, Sb, and As), increases which can result in the occurrence of the so-called slime dendrites on cathodes, thus debasing the quality of the product. As a normal suspension content in electrolyte, the contents in the order of several mg/l are considered, usually not more than 5-6 mg/l, while contents exceeding 20 mg/l are referred to as an alarm state. Suspensions present in an electrolyte are usually a light fraction of anodic slime whose main components are metallic silver, lead sulphate, complexes of antimony and arsenic, and sometimes even calcium sulphate; they are mainly composed of very small particles of lead sulphate and antimony complexes. Sometimes - - in case of alarm states - - the so-called ""wandering slims"" can be observed, which are composed mainly of arsenic and antimony complexes. These ""wandering slims"" make a serious technological disturbance. In normal operating conditions of the electrorefining plant, an increase of suspension content can result from such reasons as crack of filter cloth in a slime filter, turbulent flow of the electrolyte, anodic passivation, aeration of the electrolyte, and improper servicing of an electrolytic tank. It is necessary to continuously control the suspensions content in electrolyte for proper electrorefining."

Jan 1, 1989

By J. K. Whyatt

Assuring ground control safety in many mining and tunneling projects depends, at least in part, on an understanding of in situ stress conditions that will be encountered. Yet it is rarely practical to conduct more than a very limited number of stress measurements. Stresses along the route of a proposed excavation are typically interpolated from available measurements, often assuming a linear variation of stress with depth (or elevation). However, projects where multiple stress measurements are conducted often report more complex variations of in situ stress, usually in apparent relationship to geologic structures. These structures often include lithologies of contrasting elastic properties. A method is proposed for estimating stresses in these cases by first back-calculating regional loads from available stress measurements and then modeling the distribution of stress throughout the rock mass. The method has been successfully applied to a set of in situ stress measurements from the Coeur d’Alene Mining District of northern Idaho, USA. Results provided new insights into district stress conditions and the distribution of rockburst hazards along mine drifts and between various mines. This success should transfer readily to suitable deep tunneling projects.

By C. D. Mishra

The recycling potential for any gainful utilisation of the waste products is dictated by their quality, especially the chemical characterisation. The methods commonly used for the analysis of fly ash e.g. wet chemical, AAS, ICP-AES etc are time consuming. The paper describes a rapid XRF-based method for routine analysis of fly ash which involves preparation of local standards, optimisation of sample preparation technique and use of a simple PC-based software to account for interelement effects. The matrix corrected values are found to agree well with the wet chemical data.

Jan 1, 1992

By A. D. Kalogeropoulos, T. N. Michalakopoulos

This paper deals with the numerical simulation of rock cutting laboratory tests using the discrete element method. The main objective is to provide a method for calibrating numerical DEM models of rock cutting in a systematic way so that realistic results are yielded. The numerical models are developed with the discrete element code Yade, which models rocks as a collection of bonded spheres. The models are calibrated in a four-step process based on the design of experiments method and optimization. The calibrated models are evaluated both quantitatively and qualitatively against actual laboratory cutting tests, in terms of the ratio of the simulated mean cutting force to the actual one and the cutting force time series, respectively. The fully calibrated models showed that the simulated cutting process matched qualitatively the actual cutting force recordings and underestimated the mean cutting force by approximately 3%. It is concluded that well-calibrated numerical simulations of the rock cutting process can provide not only a better understanding of the process itself but also quantitative data regarding the cutting force and energy requirements.

Dec 28, 2022

By John W. Bell

A FEW years ago the writer had the privilege of examining a piece of apparatus in the Forest Products Laboratory of McGill University which engaged his attention. The Forest Products Laboratory has an excellent installation of wood-grinding machinery, and the officials in charge were naturally interested in acquiring means for estimating the fineness of various wood grinds. The problem was referred to Dr. O. Maass ( l ). Dr. Maass designed a piece of equipment to measure the surface of a wood grind and derived an equation applicable in calculating its value, as follows : After inspecting this apparatus, the writer was disposed to believe that, with some modifications, Dr. Maass' ingenious idea might be utilized to measure the surface of mineral powders. Obviously, some difficulty would be found in maintaining an even suspension of mineral particles of high specific gravity compared to that of wood pulp. The writer will not bore the reader with a description of the various machines built and tried out. Some of these were satisfactory enough to obtain some valuable information, and to bring to light a difficulty which threatened to bring the investigation to an abrupt ending. For the mineral quartz, it was found that Kin the formula above stated was not a constant.

Jan 1, 1944

By Briggs H.

The illustrated device uses the principle that the lamp can burn with a smaller inlet passage if the air is rich in O2 and that the size of the inlet can be made a measure of O2 percentage. The appliance consists of an adjustable cut-off applied to the inlet of a safety lamp of the improved Gray pattern. Precision is assessed from tables of calibration tests in air contaminated with blackdamp of CO2:N2 composiiton 1:5.67 at temperatures of 40-70 degrees F. Results are presented of tests at St. Germain's colliery, of tests in air contaminated with pure CO2 and of air contaminated with both blackdamp and inflammable gas; published ranges in mine-air blackdamp compositions are tabulated as ranging from 90.06% N2 to 43.90% CO2. When firedamp is present at 1.25-2.5%, its content must first be estimated by means of the cap and a fifth of its value then deducted from the lamp reading for O2

Dec 1, 1916

By Philip W. Johnson, Charles D. Haynes

"Abstract - The coalbed methane industry has become a prominent source of domestic natural gas, with its technology having evolved over the past 30 years. Whereas most coalbed methane wells are able to produce without major interruption over their economic lives, wells operating in an underground mining area are subject to being mined-through. If mine-through occurs, the productivity of the well is at least compromised, and may even be terminated. The economic consequences of mine-through may be relatively simple if mineral ownership and extraction rights are common for the coal and coalbed methane. However, if ownership is not common and a superior coalbed methane lease exists, the coalbed methane ownership must be compensated for lost production caused by mine-through. This paper presents a relatively straightforward method to calculate the value of the lost production caused by the mine-through. It uses technology from mining and petroleum engineering, as well as simple economics and financial management techniques.Sommaire - L’industrie de la production de méthane houiller est aujourd’hui une source importante de gaz naturel domestique dont la technologie a considérablement évolué au cours des trente dernières années. Bien que la plupart des puits de méthane houiller peuvent être mis en production durant l’entièreté de leur vie utile sans interruption majeure, les puits implantés en périphérie de mines souterraines sont sujets à être recoupés par ces dernières. Un tel recoupement peut compromettre, ou même mettre fin à la production d'un tel puits. Les conséquences économiques du recoupement par une mine souterraine peuvent être relativement simples quand les droits miniers et les droits d’extraction sont détenus en commun pour le charbon et le méthane houiller. Si toutefois ces droits ne sont pas détenus en commun et qu’il y a eu octroi d’une concession de méthane houiller de qualité supérieure, les détenteurs du méthane houiller doivent être compensés pour la perte de production résultant du recoupement par une mine souterraine. Cet article présente une méthode relativement simple pour calculer la valeur des pertes de production suite au recoupement par une mine souterraine. Cette méthode fait appel aux technologies de l’industrie minière et pétrolière ainsi qu’à des techniques simples d’économie et de gestion financière."

Jan 1, 2006

By Frank F. Aplan, Edward C. Dowling, Richard R. Klimpel

Several series of time-instantaneous samples were taken on a cell-by-cell basis in the flotation circuits of the Pinto Valley, AZ copper concentrator. These samples were used to calculate mass flow rates throughout the circuit using standard, deterministic metallurgical formulae and also a variety of more statistically oriented techniques. Because of the magnitude of the errors associated with the individual cell samples, the standard metallurgical formulae are not sufficient. Consequently, it is necessary to utilize some type of statistical technique capable of distributing the individual cell sample errors in such a way so as to provide a physically useful mass balance for the entire circuit. This paper explores the use of linear and least squares programming as one technique to providing credible mass balance calculations. For linear programming this was done by distributing the error using two different criteria involving specific linear programming formulations: least absolute sum minimization (which assumes an underlying double exponential distribution error structure) and minimax (which minimizes the maximum absolute deviation and assumes an underlying rectangular distribution error structure). These results were then compared to the more standard formulation of minimization of the sum of least squares which assumes an underlying normal distribution error structure. The least absolute sum approach was particularly attractive due to its simplicity of use and the ease in which repetitive calculations can be done with no user intervention.

Jan 1, 1990

By T. W. Wlodek

Abstract A simplified method of fatigue testing full-dimension, as-rolled, non-machined, mining drill rods is very desirable to the mining and steel industries in the classification of steels available on the market, through the evaluation of their fatigue limits and endurance strength. In this paper one such simplified method is presented and analysed. The preliminary results show many advantages of this method, such as simplicity and constancy of loading. The type of loading applied more or less simulates the working conditions of the outside fibres of the mining drill rods. Additional advantages of this method are the possibility of testing under dry and wet conditions, its quietness, and its ease of operation. Specially designed adaptors and specially applied techniques of fixing and preparation of samples make possible the breaking of a quarter-octagon hollow drill rod in a free length of uniform section in a rotating beam machine. At first glance this seems to be difficult to achieve. In the first preliminary test three different lots of steel were used. S-N curves under dry conditions were determined. These curves permit of the classification of the drill rods on the basis of the fatigue limit and endurance strength of the outside fibres of the metal, which is dependent to a greater or lesser extent on the effect of surface decarburization, surface imperfections, and special surface treatments. The average slope of the S.N curve in the damaging zone, and the fatigue limit in equal tension-compression rotating bending, were determined for SAE 1080, chromium-molybdenum and nickel-chromium-molybdenum mining drill rods. INTRODUCTION The fatigue limit and the endurance strength of mining drill rod steels are of interest to both the mining and the steel industries. These industries are greatly interested in lowering the cost of mining by reducing the consumption of these steels per ton of ore exploited.

Jan 1, 1951

By Robert A. Cortese, Lon D. Santis

This paper describes a method of measuring continuous detonation rates using standard electronic parts, coaxial cable, and a data recorder such as an oscilloscope. Three six-volt lantern batteries or a DC supply power the electronics that cost less than $100. Included are means of showing low battery power, finding the proper voltage scale on the recorder, and test triggering of the recorder. Modifications of the circuit allow for use in the field to measure charges up to 57 meters and for use in the laboratory to measure short charges. The high resistance (8.709 Ohms/meter) coaxial cable is a mass produced, special order cable available commercially. The driver circuit delivers a constant current to the cable, and as the detonation shorts the cable, a change in circuit voltage occurs. Recording this voltage change over time allows determination of detonation rate resolute to the sampling speed of the recorder. An oscilloscope or personal computer with a data acquisition card can easily record the data for analysis. Detonation rates recorded in the laboratory using the coaxial cable and the Bureau of Mines's traditional probes are presented. The system accurately measured the detonation rates of ANFO confined in steel pipe, detonating cord, and unconfined high explosive.

Jan 1, 1996

By J. A. H. Church

My paper entitled "Spontaneous Combustion of Coal in Mines;" read some time ago before this Branch, consisted chiefly of extracts from the British Blue Book dealing with the same subject. Fallowing my article, Mr. de Hart submitted a paper dealing with the application of existing methods of working in Alberta to the conditions obtaining in our bituminous mines. I must confess that the perusal of his paper did not enlighten me either as to what might be a method preferable to those in vogue or even to Mr. de Hart's ideas on the subject, nor did the ensuing discussion of Mr. de Hart's paper afford any great clarification of the problem. The outstanding idea, supplied by Mr. Dick, seemed to be that a retreating system could not be employed, as no operator would feel justified in risking the capital necessary to driving to the boundary when the whole might be jeopardized by a fire breaking out in the waste. That a fire would break out in the waste of any adit-level working seemed to be conceded by all, and with that I agree; that an ignited waste could not be safely sealed off by means of stoppings in a coal barrier-pillar appeared to be the contention of Mr. Dick, or perhaps he attributed this idea to the Mines Branch-anyhow, it seemed to loom over the discussion like a miasma. With this aspect I do not agree, as ? reliance is placed wholly on stoppings in the thick seams of South Staffordshire, even in fiery mines. In the thick coal at Exhall colliery, or elsewhere in Warwickshire (34 feet 4 inches) containing hard and soft coal in layers, a fire assuming serious proportions is sealed off. By the way, retreating long-wall methods are employed. In Fifeshire, sealing off fire in the Dysart mine has been accomplished successfully. The seam averages about 20 feet thick.

Jan 1, 1924

By Amir Karami, Derek Kinakin, Catherine Schmid

The Mathews Stability Graph Method (Mathews et al., 1981) has been used by geotechnical engineers and mining engineers for over 40 years to estimate the performance of longhole open stopes in underground mines. The method is widely used and accepted by practitioners because it is relatively simple to use and can be modified to account for observed performance at a given mining operation. While the simplicity of the method increases its accessibility and utilization, it does not intrinsically account for variability in the various inputs, including rockmass quality, structure, stress, or stope orientation. This paper presents a modified approach for the Stability Graph Method, whereby Python scripts are used to process available geotechnical datasets and combine them with actual or potential distributions of Q parameters, induced stress, structure, or stope orientation. The resulting Stability Numbers are then plotted as a contoured distribution on the Stability Graph.  This allows the practitioner to quickly incorporate large datasets into the analyses, while also developing an understanding of how variability in rockmass quality, induced stress, structure, or stope orientation could affect stope stability.  This paper presents the fundamental principles of the Stability Graph method, the Python script and workflow, and two case studies to demonstrate the utility of the tool.

Oct 1, 2024

By Jordan Scheremeta, Sudhir Tripathi, Mahfooz Alam

Injection leaching can improve leach ore recovery by displacing retained solutions, enhancing ore wetting, and creating new flow paths for efficient solution collection. This paper presents a method to assess the feasibility of injection leaching in finer-grained, potentially contractive ores (from a geotechnical slope stability perspective) using limit-equilibrium (LE) slope stability analysis and static deformation modeling. LE models were carried out by assigning peak undrained strengths to the heap leach ore material, and modeling the excess pore pressure induced during the injection leaching process using spatial function in Slope/W. Static deformation modeling was completed using the software Fast Lagrangian Analysis of Continua (FLAC) to evaluate the potential for deformations due to the excess pore pressures induced during the injection leaching process. A constitutive model was used to represent the geomechanical behavior of the leached ore material, which enables the prediction of deformation and potential stability issues resulting from excess pore-pressure induced during injection leaching process. The two-pronged approach provides a practical method to assess geotechnical impacts and guide data-driven decisions in injection leaching operations.

Feb 22, 2026

By Kugan Kuganathan

Pastefill is being increasingly considered as a viable fill by underground mine operators. Mount Isa Mines (MIM) chose pastefill to fill their deep copper mine (>1200m deep) after considering various other types of fill in the feasibility studies. During the development of pastefill at Mount Isa (1996 - to date) the Mining research of MIM developed two cost effective, reliable and simple testing methods and used. With the help of these tests the behaviour of MIM pastefill was studied thoroughly and the most optimum mix design for pastefill was achieved. The MIM's pastefill plant, the largest in the world with a capacity of 450rph continuous pastefill production and unique in the process used, has been built (1998-1999) to produce pastefill to the specification developed through extensive testing program using the Flow Cone and Mini Slump cone Tests. As a consequence, it is now possible to select the optimum pastefill mix designs and predict their flow behaviour in bore holes or pipe lines using only these tests. MIM Blended Tailings Pastefill (BTPF) uses 3-3.55 Portland cement and the cost of pastefill is much lower than the total tailings pastefill, that could be produced from the same tailings.

Jan 1, 2001

By Axel Preusse

In the Ruhr district, pipelines of public and private infrastructure are made of cast iron, steel, plastics or asbestos cement. Legal framework, technical regulations and normative standards define which pipeline materials may be used for the transmission of which kind of product Thus ground water endangering substances as well as inflammable or explosive products are only permitted to be transported in steel-made pipelines. Ground movements, due to underground mining, are transmitted via the soil to the pipeline and cause extra strain within the steel pipeline (Figure I). In extreme cases durable deformations may occur, characterized by an oval profile or a fold in the pipeline. To avoid or minimize damages, balancing elements are installed into the concerning pipeline, reducing the ground movements which might otherwise cause damages. [ ] The distance between the balancing elements is not arranged by random, but by a calculation method developed in a doctoral thesis [Spielberg, 1999] which represents the actual available standards in science and technology. The calculation method considers all relevant influencing variables with impacts on the movement behaviour of steel pipelines in mining areas, such as technical data of pipeline design (e.g. steel grade, wall thickness, weld seam quality), soil conditions (e.g. friction), laying depths of the pipelines, the operational impacts (working pressure, corrosion) and the mining impacts. As of today, mining has caused no damages to steel pipelines protected by this calculation method.

Jan 1, 2004

By Kanchan Mondal, Anthony (Sam) Spearing, Gopi Bylapudi, Joseph C. Hirschi

"The U.S. coal mining industry uses about 100 million rock anchors per year. ASTM F432 specifies the requirements for rock bolts on mines but does not address corrosion. Corrosion has been found to be an issue in Australian coal mines (Hebblewhite et al., 2003 and Villaescusa et al., 2007), where the problem has been researched and stress corrosion has been found to be a significant cause of rock falls. Corrosion is also a major concern for underground civil construction in the U.S. but has not been considered an issue or adequately researched in coal mines. Underground conditions are conducive to corrosion mainly because of water quality and humid conditions. There is a perception, however, that when bolts are fully grouted, adequate corrosion protection is offered to the steel. Research has shown that this is not necessarily the case, due to the formation of micro-cracks as the resin sets and with subsequent rock movement shearing the resin column. This paper outlines a method to determine the corrosion potential of bolts used in long-term excavations and suggests ways to mitigate such effects based on research conducted in the lab and on three mines.INTRODUCTIONResearch on rock anchor corrosion in mines has not been extensive enough in the past, and the effects of several factors in the mine atmosphere and waters are not clearly understood. One probable reason for this lack of research may be attributed to the time required for gathering meaningful data, which makes the study of corrosion quite challenging.Almost all the rock bolts used in coal mines are made from steel, hence the corrosion of iron needs to be understood. The corrosion of iron is an electrochemical phenomenon with the overall equation"

Jan 1, 2010

By A. Hartikainen, M. Rinne, F. Johansson, R. Yorke, L K. T. Uotinen, E. Korpi, J. Antikainen

"In order to determine the in-situ shear strength of rock joints, large scale testing is required. However, this is both expensive and difficult to execute. One possible method to overcome this may be to use photogrammetry to capture large joint surface roughness in-situ and downscale it to replica samples, which could be sheared in laboratory. In this paper, as a first part in such a method, a technique to digitize surface roughness and to produce replica samples for laboratory shear testing from a larger joint sample are presented. First, a thin granitic rock slice with dimensions of 1.75 m x 0.95 m of granitic intact rock was chosen for the study. The joint surface is fresh and created through tensile induced splitting. The large joint sample is digitized using photogrammetry. Then, one fullscale 1.7 m x 0.6 m geometry is cropped from the digitized joint geometry and then subsamples at 10x, 7.5x, 5x, 2.5x and 1x scales. All sub-geometries are scaled down digitally to produce 0.17 m by 0.06 m geometries. The geometries are used to make casting molds both positive and negative to produce samples with perfect matedness. The casting molds are 3D printed in polylactic acid plastic and C60/75 concrete is cast to produce a replica series. In addition to the creation of this replica series, two pilot replicas are also tested using a portable shear box with a 0.5 MPa normal pressure. The results from the pilot rounds are presented and discussed. Finally, suggestions for future research are given.INTRODUCTIONThe rock mass strength is in many cases controlled by the shear strength of rock joints, especially in hard rock masses. The shear strength is generally found to be influenced by several parameters, including normal stress, the basic friction angle of the intact rock, the joint surface roughness, the joint wall compressive strength, the matedness of the joint surfaces, weathering and infilling. The effect of these parameters for the joint shear strength can be evaluated in laboratory scale. However, the shear strength is known to be scale dependent. How the scale effect influences the strength is not fully understood. The general consensus is that the shear strength decreases when sample size increases, i.e. a negative scale effect (Bandis et al. 1981). However, some studies have presented contradictory results, see for example Hencher et al. (1993) and Kutter & Otto (1990). The extent and nature of the scale effect is still a question of debate (Tatone & Grasselli 2013)."

Jan 1, 2015

By Jon C. Volkwein

The U. S. Bureau of Mines investigated a method of using inert gas to prevent the formation of explosive gas mixtures in auger highwall mining of coal. A combination of gasoline and diesel engine exhaust gases was introduced into the auger drill hole using a short section of pipe located at the collar. Gas samples were taken and analyzed on site with infrared detectors for oxygen, carbon dioxide, methane, and carbon monoxide. Evacuated bottle samples were also taken and analyzed by gas chromatography at the Pittsburgh Research Center. These gas results were analyzed for explosibility. Personal exposure to carbon monoxide was also monitored. The highest methane level observed was 9.55 pct. The Inert gas levels, (carbon dioxide and nitrogen) were sufficiently high to prevent any ignition of the methane. Results showed that for all conditions during mining, gas concentrations were non-explosive. The maximum personal time weighted average sample for carbon monoxide was 20 ppm. This system provides a safe, inexpensive, simple method for preventing explosions during auger mining. INTRODUCTION The auger highwall mining method is an effective method to recover coal from a reserve when removal of the overburden by surface mining equipment becomes uneconomical. In this method of mining, a horizontal auger enters the coal seam from the surface mine bench under the highwall and the coal is drilled in a series of parallel holes. Historically, coal mined from the surface is relatively shallow, and over time, methane associated with the coal has dissipated through the surface. In most circumstances, little methane has been found associated with auger mining. However, mining technology has enabled surface mining of deeper reserves of coal. Furthermore, environmental constraints have forced the highwall extraction method to be used to remove coal under wetlands, further increasing the chances of encountering methane. Recently incidents of methane explosions at a few auger mining operations have resulted in injuries and increased testing for methane at the collars of auger holes. The fuel source of the reported explosions was not necessarily limited to methane, but may also have involved coal dust. The Mine Safety and Health Administration (MSHA) met with the Bureau to discuss what technology might be available to enable the safe resumption of mining. The discussion included the difficulty of ventilating through the solid shafts of the augers, that steel bits probably created the ignition source, and that perhaps inerting the holes with low oxygen and high carbon dioxide concentrations from the machine's diesel exhaust was a potential solution. Considering the ventilation aspects of the problem, it was not clear If ventilation could be reliably established. If some degree of ventilation to the front of the mining head is achieved, it may combine with methane to bring the hole atmosphere from a rich, nonexplosive mixture to an explosive mixture. Furthermore, it may not prevent a dust explosion in such a mining configuration. Lack of access through the shafts of auger type mining machines further limits the ability to add water or air to cool bits to prevent an ignition source from developing. Either of these approaches would also be expensive. The process of mining coal In an inert atmosphere has been considered in the past, but to our knowledge, never implemented (Department of Interior, 1970). Clearly, implementation in underground mining would be more complicated. On a mine bench open to the atmosphere, however, adding inert gas to the mining head could provide a quick, feasible method to prevent explosions at auger highwall mining operations. Also the problem of how to move the inert gas to the cutting head of the machine had to be considered. Preventing explosions on auger mining machines using inert gas requires three primary considerations: first is the source of inert gas; second, placing the inert gas at the cutter head; and third, monitoring the hole atmosphere. Any gas source having an effective inert gas concentration of 34 volume pct or greater will prevent methane from Igniting (Zabetakis, 1965). Sources of inert gas considered for this application included liquid nitrogen, modified shipboard inert gas generators (for hydrocarbon shipping and transfer), jet turbine engine (Paczkowski, et. al., 1982), the auger's diesel engine and a gasoline engine. Operation cost, purchase cost and availability limited our testing to the diesel and gasoline engines. This work tested each engine, separately and combined. To ensure effectiveness, both company and enforcement personnel need to know how to monitor the condition of the inerted hole. Measurements at depth inside the hole are possible by remote sampling through rigid tubing, but this method is Impractical for routine monitoring. Continuous monitoring of the exhaust gas stream is an alternative. The U. S. Bureau of Mines evaluated an inert gas system at an auger mining operation at a surface mine near Owensboro, KY. Coal was mined from the Number 9 Coalbed in Henderson Co. KY. Tests were conducted in January and March of 1992.

Jan 1, 1993