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By M. A. Anderson

After mope than fifty years Of Continuous operation, the Nevada Mines Division found itself faced with declining ore headings and metallurgical efficiencies, and rising operating costs. Although intensive process improvement programs had been initiated at various times in its history, insufficient return on investment served to defeat plans for any major changes in circuitry. The advent of practical large-volume flotation machinery, along with the success of long-range exploration drilling programs, complimented the results of an intensive program of research and development. The result of laboratory testing, pilot plant testing, and full scale plant tests was the proven indication of metallurgical improvement given increased flotation residence time and the effect of mechanical agitation in flotation.

Jan 1, 1975

By A. B. Van Cleave

It was first pointed out in 1934 by Gaudin (1) that the "oily" oxidation product of the alkyl xanthate ion-dialkyl dixanthogen, might be the "active flotation agent" on sulphide minerals and there have been a number of studies supporting this hypothesis (2,3,4). However, all of these studies have relied largely on electrochemical evidence. Galena, being a semiconductor, may be used as an electrode in aqueous solution and it was shown that ethyl xanthate ion does not adsorb with charge transfer in the absence of oxygen unless the potential of the galena surface is above -100 mv versus the hydrogen electrode (H2 E). At this potential the galena surface is slightly hydrophobic but does not become strongly hydrophobic unless the potential is raised to 210 mv (versus H2 E). It is suggested that, at potentials of 210 mv or above, ethyl xanthate ions are oxidised to diethyl dixanthogen and that chemisorption of oxygen raises the galena surface to this potential. On the other hand, examination of xanthated galena surfaces by chemical or physical methods (5,6,7) has failed to detect dixanthogen and it has been shown that dixanthogen may be reduced to lead xanthate at a lead sulphide surface. (5,8)*. It was suggested by Leja and Poling (5) that, as dixanthogen is a neutral molecule it would not be repelled by the negatively charged galena surface so that the adsorption was envisaged as taking place through homogeneous oxidation of the xanthate ion to dixanthogen, adsorption and reduction of the dixanthogen to a chemisorbed lead xanthate species. In paper I of this series (11) it was shown that the thiosulphate ion, which forms on galena surfaces in slurries at the pH normally used for flotation, will reduce dixanthogen. There is, therefore, the impasse that the electrochemical evidence requires that xanthate ion be oxidised to dixanthogen at the galena surface while the chemical evidence shows that the reverse reaction takes place.

Jan 1, 1973

By D. S. Choi

The paper presents the result of analysis of subsidence data collected from coal mines operating in Northern West Virginia and Southwestern Pennsylvania. A particular emphasis la placed on the measurement of the angle of draw for the protection of surface structures.

Jan 1, 1986

By John E. Butler

Due to the present economic climate, particularly the depressed copper market, conventional mining, milling and flotation operations have been suspended. However, due to the apparent amenability of the ore to 'natural' leaching, investigations have been initiated into establishing an in-situ leaching operation as a preferred alternative to conventional operations. Cement copper production has been developed in parallel with the leaching trials and is presently at the rate of approximately 3,000 tonnes of contained copper per annum. Testwork includes controlled leaching of ore both in the laboratory and in-situ following blasting.

Jan 1, 1981

By David G. Chedgy

It is probably safe to assume that most operating personnel working in coal preparation plants are exposed to noise levels in excess of the MESA Noise Regulations unless they are wearing ear protection. The excessive in-plant noise levels have received little attention until recently. Noise control has not been an important design consideration in existing coal preparation plants. This situation is rapidly changing as plant owners and plant operators are requiring that new coal preparation plants be designed to provide a safe noise environment for operating personnel and also meet MESA Noise Regulations. This paper discusses the noise exposure hazards, noise generating mechanisms, and some possible noise control solutions for typical coal preparation plants.

Jan 1, 1975

By R. R. Klimpel

In this first report, we present a preliminary model based on the simple binary sys¬tem of component B occluded in a matrix of compo¬nent A, for the case where fracture planes can pass with equal ease through A or B. By assuming that the volume fraction of B in AB in the overall matrix is small, which is realistic for many prac¬tical systems, the problem is reduced to a one-¬dimensional treatment which is conceptually easy to understand. The model was programmed for Monte Carlo simulation, and an analytical solution also developed based on the assumption of zero or one mineral fragment per particle. Preliminary ex¬perimental confirmation using a synthetic "mineral¬-host matrix" of known amount and size of B in AB is presented.

Jan 1, 1983

By C. D. Iles

The philosophy used in designing and writing Duval's Sierrita mine design programs is discussed. The Sierrita Incremental Design system (SID) is reviewed with emphasis on Pushback Extraction Design (PED). FED is a method by which predesigned pushbacks are sequenced into an optimized period-by-period design that can be directly utilized in daily operations as well as intermediate and long range planning. The advantage of SID is that it provides a practical, mineable open pit design. PED is not a theoretical design system; it is in operational use.

Jan 1, 1981

By David Green

Outokumpu introduced Flash Flotation in the early l 980's for the recovery of floatable material from grinding circuits. The devel­opment came as a result of surveys made in the company's own concentrators, initiated to find ways of improving recoveries with increasingly complex and lower grade feeds. Hydrocyclones make a separation on the basis of particle size and mass, resulting in the "over collection" of fine grained, high S.G. particles. Minerals such as pentlandite, galena and gold are easily caught in the circulating load around the mill until they are overground and rendered more difficult to collect. In the case of downstream flotation this may simply involve a reduction in the rate coefficient due to the finer sizes or, as is commonly experienced in gold operations, the mate­rial is slimed and coated onto non-floating gangue and is eventual­ly lost to tails. At the time of Outokumpu's plant surveys, it was a relatively common practice, particularly in Pb operations, to have a flotation section fed directly from the secondaiy mill discharge. This is called unit flotation and certainly provided an improvement but very high wear rates and frequent sanding out of the cells was com­mon. The development of the SkimAir® Flash Flotation cell by Outokumpu provided a subtle but quite dramatic departure from the unit cell approach. Now it was possible to feed cyclone under­flow into a special purpose-built flotation machine and recover a final grade concentrate from the grinding circuit (see Figure I). The cell was designed to handle the coarse circulating load, per­mitting the coarsest fractions to be short-circuited back to the mill and retain the floatable sizes for a nominal time of 30 seconds to several minutes. The short retention time was possible largely due to the relatively clean content of the cyclone underflow stream. This means that the fine gangue is not present except in small quan­tities and most of the fine to mid-size paiiicles are higher S.G. and more likely represent valuable mineral. This condition provides for very high rate coefficients. Typical operations would recover 40-60% of the valuable mineral in this fashion and improvements in recovery were typically from 2-5% for sulphide minerals to the 5-20% range for gold and silver.

Jan 1, 1998

By F. A. Gates, Bengt Samuelson, James E. Edmunds, S. McCune, Neil Hario, Norman L. Weiss, J. M. Bertram, A. M. Cavaliere, F. M. Jr. Stephens, D. D. Chiang

DESIGN BASIS Size of Project The size of the project generally is expressed in tons of ore milled per day, but company policies differ widely in this respect with a few adhering to this set figure but many expecting the design capacity to be liberal. To some companies that follow the latter practice a 25,000-tpd milling plant that cannot reach 30,000 or 35,000 tpd with¬out major additions is a disappointment. The engineers responsible for planning the operation must know what is intended. Other projects are sized according to the tonnage of product rather than of mine ore; examples are iron ore mills that express size in terms of tons per year of pellets or high-grade concentrates, heavy minerals plants in tons per year of rutile or ilmenite, cement plants in barrels per year, potash in tons per year of muriate, phosphate in tons per year of washer rock and flotation concentrate, and others. However, tons per day milling rate is also the common denominator in most of those instances because the major equipment has to be selected on that criterion. Equally important are operating schedules, particularly the number of operating days per week and shifts per day for ore delivery, crushing, grinding, and processing. Hours per shift also are an impor¬tant factor in design, particularly in the case of ore delivery, and the lost time expected for maintenance, cleanup, inspection, preventive maintenance shutdowns, and many others, has to be taken into ac¬count. Frequently, the ore deposit will vary in hardness and grade to such a degree that the capacity of the mill will vary widely from year to year and cannot be expressed as a fixed figure, but rather as an average over a calculable period, such as ten years, or as a specific daily capacity over the first two years, next four years, etc., or some similar projection. A good example of this kind of situation is Asarco's Mission operation 15 miles southwest of Tucson, Ariz., described by Weiss and Vincent,1 where the major rock types differed widely in hardness, so that a change in the proportion of these types in the mill feed could have imposed difficult design problems if a constant tonnage rate had been expected. Expected Life The life of a mining operation depends upon the size of the deposit and the rate of mining. The latter is a policy decision based on an economic analysis2 which, in turn, is based on many factors. From the point of view of mill planning, the estimated life of the operation determines many criteria and affects many decisions on strength of structures and degree of protection of men and materials, the quality of the materials-handling and process machinery, the type of delivery systems for ore to the mill and concentrates to the smelter or market, environmental considerations, and many others. Costly installations that may be excellent investments for a 20-year project may be waste¬ful for a 5-year operation. An example is the comparison of different methods of crushing or grinding. Take single- vs. multiple-stage grind¬ing-the former may be found more economical for a 5-year operation, and the other, more costly method more economical for a 15- to 20-year life. Mining and Ore Transportation The mining method and the means of transportation of the ore to the first crushing stage have far-reaching effects upon mill design, particularly the crushing sections, and upon site selection (see Chapter 2, Situation of Mill). Consequently, mill design cannot advance until the mine planning has reached at least tentative decisions concerning these important matters, or reach the final stages until much is known about the ore that may be expected month by month and year by year for a reasonable period. Methods of ore transportation used in over 90 underground and open-pit mines are listed briefly in Section 32 (Table 1), and more detail is found later in this section and in Section 10, Storage and Transport. These methods include surface railroad, trucks, conveyors, surface skips, underground skips, and tramming from hoist or tunnel. In surface mining, transportation of the uncrushed ore by rail or truck to the primary crushing plant-conveniently located for lowest overall cost-is most common. Rail transportation provides low perton-mile costs. Truck transportation, though in some cases less eco¬nomical, is more flexible and more adaptable to mountainous terrain; furthermore, recent improvements have reduced the operating costs to keep trucks competitive on low-grade ores. The design of the primary crushing plant will be essentially the same whether delivery is by truck or rail, but the location of the plant is subject to more rigid restrictions in the latter case because of grade limitations, and these in turn may adversely affect the selec¬tion of the mill site as well. The influence of truck haulage in reducing the stripping ratio in open-pit mines is an added reason for predicting a minor role for rail transportation in the future. The use of conveyors to transport ore from the pit to the milling plant involves (in most cases) crushing the ore to a reasonable size (10 to 20 in. maximum with today's crushers) before conveying. The Twin Buttes system3 which went into operation in late 1969, utilizes primary crushers (two) and conveyors to receive rock waste, ore, and low-grade stockpile material from trucks, then transport them to the appropriate destination. Flexibility is dependent upon the porta¬bility of the primary crushers. In the face of this development, inclined skip hoists are no longer important to the mill designers. Underground mines generally pose fewer problems of design; these will be discussed in Chapter 2. Character of Deposit The degree of homogeneity of the ore in various parts of the mine is an important factor. The characteristics that concern the millman (approximately in order of importance) are: hardness, assay, metallurgical response, size distribution,4 and moisture content. Under metallurgical response may be included degree of oxidation, mineral¬ogy, clay content, and reagent consumption, as well as process re¬sponse in general. Short-term variations-hour-to-hour, day-to-day, or even week-to-week-make operation of the plant more difficult, require greater flexibility of process design and, perhaps, also require more instrumentation and manpower. Uniformity of the ore in the characteristics noted previously and

Jan 1, 1985

In recent years there have been significant advances in computing power and programming technology, enabling more complicated processes to be modelled in greater detail. This has led to increased understanding of these complex processes, including mineral processing circuits. JKSimMet was developed over the past 35 years, incorporating outcomes from various research projects conducted at the Julius Kruttschnitt Mineral Research Centre (JKMRC), part of the Sustainable Minerals Institute at The University of Queensland. It is a steady-state software program that allows the user to mass balance, model fit and simulate crushing, grinding and classification circuits. JKSimFloat has been developed over the past 20 years, incorporating outcomes from the JKMRC and other institutions around the world through the AMIRA International P9 project ? The Optimisation and Simulation of Mineral Processing Circuits. It is also a steady-state software program, similar to JKSimMet, but is focused on analysing flotation circuits. Both programs have been used in a wide variety of optimisation and design projects over many years. This paper provides an overview of both JKSimMet and JKSimFloat and examples of how both programs have been used to better understand, develop and optimise mineral processing circuits around the world.

Feb 27, 2013

By N. Dhawan

With ever-increasing demand, low-grade complex ores and strict environmental regulations, metallurgical processing options have become limited. Heap leach technology provides several benefits to combat these issues including low cost, flexibility and diversified process conditions. This paper provides an overview of heap leaching technology with specific emphasis on chemical aspects. Hence, influencing factors, advantages and disadvantages, testing approaches and some fundamentals will be discussed. Finally, heap leach technology for different ore types (copper, nickel, uranium and gold) will be briefly compared.

Jan 1, 2012

By J. P. Mogan

As diesel exhaust contains a mixture of several materials of varying degrees of toxicity which interact to a currently undefined extent in their effects on health, the prescription of a quantity of ventilating air to dilute the exhaust to a safe level is necessarily, to some extent, arbitrary. Some jurisdictions rely on the application of a safety factor to a single "governing" component (usually nitric oxide for most modern diesel engines). Others specify that the total of the "adjusted" concentrations (actual concentration divided by maximum permissible concentration) not exceed "one" for related constituents (irritant gases for example). Neither of these approaches, however, takes account of the interaction of the gaseous constituents with diesel soot (although many studies verify such interactions), nor do they provide any economic incentive for reducing pollutants which are not included in the criteria for ventilation. Because of these factors, the Canadian Department of Energy, Mines and Resources awarded a contract for a study of the current literature bearing on the "Health Implications of Exposure of Underground Mine Workers to Diesel Exhaust Emissions" (1). One goal of this study was the formulation of a new ventilation criterion. The contractors, Ian W. French and Associates, studied more than 1500 literature citations which they interpreted in a 350 page report and summarized in a companion 1600 page annotated bibliography. In the recommendations section of the report, they defined a new 'health effects index' or ventilation criterion "V" as follows: [ ] The contractors consider that a "V" between 3.0 and 4.0 poses a moderate threat to health which could be alleviated by the use of protective equipment such-as filter helmets. If V exceeds 4.0, an increase in ventilation quantity to reduce the pollutant levels is indicated. In the case of mines with a quartz content in excess of 20% in the ore, the contractor recommends that a term. [Quartz]/ [Threshold Limit Value Quartz] be added to the equation, and [2.0 = ?v? = 30] be the range justifying protective equipment, with added ventilation if "v" exceeds 3. A degree of pragmatism was inherent in the development of the equation, as the contractors recognized that the components of the index, if it is to be useful in the near term, should be within current routine underground environmental assay capability.

Jan 1, 1980

INTRODUCTION The role of comminution and related unit operations in mineral processing applications is of major importance. This is especially true in terms of operating costs, since these unit operations represent the major fraction of the overall processing costs. Due to the broad scope and the large number of unit processes and machines involved, Section 3 is divided into five subsections. In Section 3A, General Aspects of Comminution, topics covered include history, terminology, particle¬size distributions, classes of comminution machines, testing and calcu¬lation procedures, and simulation and modeling. In Section 3B, Crush¬ing. the subjects include jaw crushers, gyratory crushers, cone crushers, hammermills, and impactors. The chapters of Section 3C, Grinding, cover aspects common to tumbling mills, ball mills, rod mills, autogenous mills, pebble mills, roller mills, fluid energy mills, other types of grinding machines, and grinding costs. Section 3D, Classification, coverage includes classification theory, hydrocyclones, and mechanical classifiers, and in Section 3E, Screening, the topics presented are classes of screens, selection of vibrating screens, sieve bend screens, and screening media. About five years after the Handbook was begun, Section 3 was unfortunately without an editor or contributors for over half of the chapters. A committee was then organized to expedite completion of the section. When this committee contacted the persons who had agreed to contribute to the section, it was found that many of them wished to resign from their earlier voluntary obligation to AIME. Therefore the committee had to find new contributors for these chap¬ters in addition to other chapters in Section 3. Since there had been no previous effort to locate contributors for Air Classification, Dust Collection, and Washing and Scrubbing, it was agreed to delete these subsections. In Section 3, equipment availability, capacity, and performance data from various manufacturers and users are provided for engineer¬ing use. Selected data from existing installations are also included. The machines that are being used in comminution, classification, and screening are described in detail, and drawings, photographs, and diagrams are presented to supplement the descriptive material. The editors thank and acknowledge the generous effort of the many contributors, some of whom were called upon five or more years after the Handbook was started. It is regretted that Fred C. Bond, who was a member of the committee formed to complete this section, was deceased before the section was finished. During the time that he worked with the committee, he contributed a great deal as an author and councilor.

Jan 1, 1985

By Jurgen H. L. Stuve, Rick E. M. Widelski

Any one who has been involved with the mining or the treatment of gold-bearing ores will agree that at the end of the total process, the quantity of gold recovered is never as much as was expected. These short falls are usually explained away by claiming that:- the original evaluation of the ore being mined was inaccurate the plant treatment process is not at its optimum theft is taking place excessive dilution occurs during mining The truth of the matter is that the gold is being "lost". During all the stages of its production, starting from the mining, right through to the smelting process, a certain amount of gold is not recovered. When a newly found gold deposit is being evaluated and the drilling data being analysed, cautionary factors are applied so that the grade is not over-stated. Once the decision has been taken to start mining and development is in progress, the exposed ore is sampled. When Again, when this data is being analysed more safety factors are used to ensure that when a piece of ground is given a block-value, the gold content is not over estimated. mining actually takes place and the ore transported to the plant for treatment, the metallurgical department also introduces certain safety factors when predicting how much gold will be recovered. This shows that the quantity of gold that one expects to finally recover is already a conservative figure. The actual quantity recovered invariably falls short of this figure, in fact, South African gold mines are satisfied when achieving a Mine Call Factor (MCF) of 85%. In the last few years a great deal of work has been carried out to investigate how this unaccounted for gold is lost and what steps can be taken to remedy the situation. Previously mined areas were re-swept using the traditional hand brush method; experiments were carried out with water jets sometimes in conjunction with compressed air and even small portable vacuum cleaners were tried. None of these methods were totally successful in recovering the left over gold but the use of vacuum systems showed the most potential. the In South Africa this has resulted in introduction of high-powered vacuum systems on the gold mines to suck-up this lost gold. These were initially introduced in the mining areas, particularly old areas, to recover any left over ore. The results proved extremely profitable, So much SO that presently, well over 120 suction systems are doing this type of work. This success encouraged plant operators to look at areas of potential loss in the plant itself. This has resulted in an increasing number of machines being supplied to gold plants mainly for clean-up applications with the bonus that the material sucked up invariably contains gold.

Jan 1, 1993

By James Stewart

"THAT WHICH WE CANNOT SEE" INTRODUCTION Often it has been stated that the greatest of all fears is fear itself - fear created only by that which we cannot see or by that which we fail to understand. The elimination or minimization of fear can best be accomplished by a thorough understanding of the underlying causes of the fear. As related to the mining industry the omnipresent fear is the dread of acquiring cancer through accumulated exposure to "that which we cannot see" - radioactivity. The understanding of the resulting exposure effects can best be obtained through acquisition of authentic data and the subsequent analysis of the same by a group of highly qualified specialists. I may point out that the nuclear industry and its source, uranium mining, greatly surpass any other industries in the degree of public awareness as to its potential environmental effects. Public out-cry has always influenced major decisions relating to how industries conduct their operations and much of their effort has resulted in many changes. The issues of health and safety in the mining industry have created considerable public concern, but intense concern to those of us directly involved within the industry. Increasing concern regarding the effect of human activity on the quality of the outer environment of air, land, and water has developed within the past two decades. There is now a growing concern about the health and safety of persons at work in their inner environment, of their working-places in industry. Hearings and investigations conducted in Canada by three appointed "Royal Commissions" have all confirmed that this concern is well justified when focused on the health and safety of workers in our mines. REGULATORY CONTROL OF URANIUM MINING OPERATIONS In Canada, the ownership of mineral resources and the control of their extraction, except for uranium, is vested under the "British North America Act" in the provinces. In all provinces the statutory basis for health and safety provisions in the mining industry can be found in the provincial "Mining Acts" with each province assuming the responsibility to produce its own "Mining Act". In particular, all of the Mining Acts in Canada place upon the management of a mine the responsibility to ensure that provincial provisions are met. The "Federal Register" and mining companies parallel this arrangement in the United States. Equivalent legal

Jan 1, 1981

By Bahne C. Cornilsen, X. Shan, David H. Carlson, John H. Johnson

The air in three U.S. underground coal mines has been sampled. The data developed show the relative importance of the various pollutants and quantify the need for control. Total respirable particulate matter (TRPM) was found to be the critical pollutant (the one requiring the most dilution for control). TRPM is considered to be "respirable coal dust" by procedures approved for sampling in U.S. underground coal mines. However, this "respirable coal dust" also includes diesel particulate matter (DPM) and particulate from other sources. While "respirable coal dust" was the critical pollutant, Raman spectroscopic analysis showed that more than half of this "respirable coal dust" was DPM. The DPM in the samples ranged from 22 to 152% of the 2 mg/m3 limit for respirable coal dust. Thus DPM was the critical pollutant using the 2 mg/m3 respirable coal dust limit. SO2 gas produced by the combustion of diesel fuel containing sulfur was the second-most critical pollutant. Characteristic curve slope (CCS) values were calculated to determine the quantitites of the various pollutants generated per unit of diesel haulage activity. Mine 1 had the highest TRPM CCS values with 29 mg/m3/%C02 for the Ramcar and 57 mg/m3/%C02 for the Haulageway location. The highest coal particulate CCS measured was 30 mg/m3/%CO2 and the highest diesel particulate CCS measured was 27 mg/m3/ %C02. Both of these high values were measured in the Mine 1 haulageway location. Calculations were made to determine the controlling CO2 concentrations (CCC) at which all other measured pollutants would be within their limits. These CCC values, which, for the mines sampled, were based on TRPM, varied from a low of about 0.07% for the Mine 1 haulageway location, to a high of about 0.15% for the Mine 2 Ramcar and haulageway locations. Weekly average CO2 concentrations measured in these mines ranged from 0.09 to 0.11 % for the Ramcar, from 0.05 to 0.09% for the haulageway, and from 0.08 to 0.11 % for the return. The concentrations of TRPM ranged from 1 and 2 mg/m3 in mine areas frequented by personnel, and to greater than 4 mg/m3 in returns where personnel are not exposed.

Jan 1, 1991

By R. Allan Wallis

In the past mining engineers have tended to rely heavily on the advice of fan manufacturers in relation to fan equipment issues. A more adequate approach was advocated in a paper [ Wallis, 1965 ] submitted to the Eighth Commonwealth Mining and Metallurgical Congress held at Surfers Paradise, Australia during 1965. At that time Mount Isa Mine were in the process of installing 3 large 6096mm (20ft) diameter vertical axial flow surface fans of an advanced type [ Mathews et al, 1967 ]. The preliminary feasibility studies prior to preparing the detailed job specification for this project were extremely thorough with the result that it was possible to include detailed information and instructions in the document, ensuring the procurement of equipment that was satisfactory in all respects. Full advantage was taken of expert services provided by external organizations during the large coordinated multi-disciplinary exercise. My continuing activity and development in relation to the technology of axial flow fans is now at a stage where all known design and performance data are accessible through CAD programs [ Wallis and Kipp, 1987 ]. The extensive calculations that were necessary in 1983 can now be carried out in a mere fraction of the time, and with complete confidence. Model testing in order to validate the fan and duct design proposals is, in my view, no longer essential. The first of the computer programs enables a mine ventilation engineer to study a large number of alternative axial fan arrangements for comparison with the centrifugal alternative. By entering the volume flow rate, fan pressure and air density, he can then with ease, and rapidly, investigate the effect of various blading arrangements, downstream evasee area ratios, and hub ratios, on fan diameter and speed, with their related attainable fan efficiencies. This procedure had its origin back in the 60s when I was advising mining companies on new fan equipment. At that time such an investigation was quite lengthy and lacked the precision and completeness that is now available within the present programs. To obtain the major benefit from these outputs, it is most important not to be disconcerted by the three very important nondimensional parameters, namely, the tip flow coefficient (LAMt), the fan pressure coefficient (Kt or Kit), and the swirl coefficient (Es and/or Ep) as these are important when assessing the various design proposals. When seeking an understanding of the role and importance of each of these variables, ample help is provided in the user manual; the design recommendations are summarized in the help screens which are readily accessible within the design programs.

Jan 1, 1989

By D. H. Carlson, J. H. Johnson, C. F. Renders

Research was conducted at a western coal mine on a wide variety of underground mining vehicles to develop testing procedures for portable instrumentation monitoring of tailpipe concentrations of diesel particulate matter (DPM), carbon dioxide (CO2) carbon monoxide (CO) and nitric oxide (NO). DPM was measured using the Emissions Measurement Appa¬ratus (EMA) developed by Michigan Technological University (MTU), and an ECOM-KL five-gas analyzer was used to measure gaseous tailpipe concentrations. Baseline concentrations of the tailpipe pollutants were measured using vehicle systems to load the engine. Concentrations before and after the water scrubber were measured and water scrubber efficiencies were estimated by calculation. Field measurement procedures were established. Heavy-duty vehicles with automatic transmissions were loaded using the torque converter to stall the engine. Consistent engine loading for light-duty vehicles with manual transmissions involved placing the transmission in neutral and adjusting the throttle to maintain a constant engine speed of 2,500 rpm. Tailpipe CO2 concentrations for Caterpillar 3306 and 3304 engines were compared to laboratory CO2 data from Caterpillar and MSHA to help determine if engine settings were correctly adjusted for operation at high altitudes. (Perfect agreement between the mine vehicle tailpipe and laboratory measurements would not be expected because of the differences in the test conditions.) DPM concentrations for light-duty vehicles ranged from about 20 to 160 mg/m3, and one light-duty vehicle averaged near 1,000 mg/m3. DPM concentrations for heavy-duty vehicles ranged from about 55 to 200 mg/m3. Widely ranging DPM concentrations for a particular vehicle type reflect the state of maintenance, fuel rate settings and different engine loading procedures as well as differences between samples collected upstream and downstream of the waterscrubber. The calculated DPM removal efficiencies due to the water scrubber varied considerably between vehicles, ranging from 13% to 51%. Calculations indicate that, under worst-case conditions, very large reductions in DPM may be required to meet the proposed DPM PEL of 0.15 mg/m3, even on vehicles with water scrubbers. In-mine ambient air studies, which are needed to actually measure the impact of the tailpipe concentrations on the work environment and the effect of mine ventilation on the dilution of DPM were not done in this first set of tests.

Jan 1, 2000

JIM ARNOLD: Questions? Comments? Gary, were you gone when I made the comments about corporate puke? [laughter] SPEAKER: You were hiding. BRUCE CAVENDER, BHP Copper: Having worked in both places, I'd like to address the issue of the plant versus the corporate office. One thing to keep in mind is that we are all very engaged with the technical elegance of a lot of the work we do. There's some tremendous technical expertise in the plant. But remember, the guys in the corporate office are only thinking in one denomination, and that's dollars. So we'll see something like a spill, and for us it's a problem. At the end of the day, what the corporation cares about is the financial impact. So we've got to be careful to couch our solutions in dollar terms. JIM ARNOLD: Rebuttal? [laughter] DOUG HALBE: Are you going to rebut this one, Andy? ANDY CARTER, Oxidor Gold Corporation: No, I don't think so. I'm not up to it. I'd just like to address a question both to our panel and the floor. Those of us who've been in this industry any length of time have seen these same mistakes made over and over again. Wherever you go, year after year, plant after plant, you see the same problems. I guarantee you, Doug, that if we have a similar conference down the road, you're going to see the same thing. My question to the panel and to the floor is what is it about our industry that stops us from learning from our mistakes? What characteristic is it that prevents us from learning and makes us repeat the same mistakes over and over again? DOUG HALBE: That's a great question. Someone said in a meeting 10 years ago that if the people who design plants designed airplanes, the planes would be falling out of the air all the time. GARY SIMMONS, Newmont Gold Company: I've got an answer to your question as to why many plants are poorly designed. It's that the minute anybody demonstrates any sort of competence in plant design, they're promoted to a higher-paying job, and the talent pool is left with people of lesser experience. Just think about it: The people who have the experience and have demonstrated it more than once have moved up. JIM ARNOLD: I think that there's another reason, too. Seven or eight years ago I was chairing a session, and I challenged everyone in the session to some day publish an article about the design mistakes they'd made because we never talk about our mistakes. I'll

Jan 1, 1998

By Ian S. Lowndes, Adrian T. J. Moll

With the current trends towards the concentration of extraction to fewer but higher production units in deeper seams, often of higher coal rank, the role of methane drainage ranges has become critical. As mines exploit seams at greater depth and more remote locations there has been a need to extend the lengths of methane drainage ranges and to increase the capacity of the surface pumps. This trend has highlighted the need to effectively simulate the operation of a methane drainage range in order to produce the optimum design and operation of a safe and economic system. This paper describes the development of such an analytical tool. The application of this computer model is illustrated by the analysis of the operation of the drainage system of an extensive mine. The paper concludes by considering the development of a [Knowlege Based or Decision Support System] which will assist the ventilation engineer in assessing the current performance of an existing drainage range and will suggest possible remedial modifications or design strategies which may improve the system performance. INTRODUCTION In order to maintain the general body methane levels to below predetermined statutory threshold limit values (TLV's) on major longwall production units there are, for a given production rate, two major complementary control devices which may be employed by the ventilation engineer. These are the combined design and application of a competent mine ventilation and methane drainage system. The methane drainage system should be designed so as to maximise the capture of the methane (firedamp) from the seam being mined and the strata gas liberated in the vicinity of the extraction zone from gas bearing seams or other reservoirs of gas situated at above or below the mining horizon. The design fresh air quantity entering the face should ensure that under normal operating conditions the methane entering in the vicinity of the face is rapidly dispersed and dilute to well below statutory TLV's. The drainage system must therefore be designed to maximise the capture of the methane gas liberated to minimise the amount of gas which enters the ventilation airstream. GAS DRAINAGE The inherent dangers of methane emissions in mine workings have been a problem of concern to mining engineers for many years. To alleviate the risk, methane drainage has been employed for some years to capture the gas before it becomes a hazard to the operation of the mine. The method of drainage employed will be dependent on a number of factors which include the longwall mining method and production cycle employed, the characteristics of the coal seam being extracted and those of the adjacent gas bearing strata. The location of the gas bearing potential of the over and underlying coal measures must be determined. The gas bearing

Jan 1, 1993