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By Gregory P. White

A number of private companies and universities have developed computer programs which can be used to simulate surface coal mining by dragline. In this paper, we briefly discuss the concept of computer simulation, compare these computer programs with each other, describe the major changes required to convert these programs to an IBM computer system, and analyze their sensitivity to inaccurate input data. We also present a summary of the required data input for each program, core storage requirements, and output results. Sources for obtaining these programs and their documentation are listed at the end of the paper.

Jan 1, 1983

By Steven M. Puricelli

In recent times, smelter off gases treatment plants have become more efficient and environmentally friendly. Where in the past it was acceptable to process only high strength, treatable furnace gases, new smelters must now treat primary gases and abate dust and SO2 from secondary gases: Processing smelter gases can present some interesting challenges. Some equipment can see varying concentrations of sulfuric acid, up to the boiling point, with a periodic table of contaminates thrown in for good measure. Treating weaker secondary gases have, in the past, been cost prohibitive. Nowadays, new technology, like physical absorption or catalytic carbon, show promise for economically treating low strength secondary gases. Modem smelters do a much better job of treating the SO2 laden gases generated by the smelting process but significant challenges remain.

Jan 1, 1998

By Z. Kokesz

The following paper presents the results of geostatistical studies of the selected hard-coal deposits in Poland. Geostatistical methods were applied mainly to the characterization of variability structure of the coal seams thicknesses as well as sulphur and ash contents in the coals. Results imply that the accuracy of coal reserves estimations can be improved if kriging method is applied. Kriging allows predicting more precisely the values of such parameters as thickness and sulphuring content.

Jan 1, 1993

By M. Z. Emad, M. U. Khan, A. ur Rehman

Stemming plugs are used as an accessory in drilling and blasting industry. They supplement conventional stemming material (i.e. drill cutting) with retention of gases in the drill hole that result in better fragmentation of rock and improved explosive utilization. Better fragmentation and explosive utilization along with economic benefits enhance efficacy of the project. Incorporation of stemming plugs in drilling and blasting industry is in exercise for about thirty years. The efficiency of stemming plugs is a key factor in their rising induction in mining industry. Different studies have been conducted on the economic viability of these plugs. The focus of this paper is the selection of stemming plugs for a bench blasting operation and the examination of blasting operation efficiency based on ergonomics and human factors. Three types of stemming plugs were evaluated during multiple blasts at a limestone quarry. Their performance was evaluated using split desktop software. Based on the human factors plastic molded stemming plug (stemming plug-1) was found to be the best among all selected plugs. INTRODUCTION Surface mines and quarrying operations heavily depend on bench blasting to achieve high volumetric targets and production demands. Use of explosives is the most economic method adopted for rock breakage and movement. Extensive research had been performed to enhance explosives energy utilization. This include changing the drilling pattern, application of air decks, stemming plugs, and use of other supportive devices. Optimization of explosive energy has been a major concern in those mines that rely on drilling and blasting operation to carry out main production. Due to continually increasing cost, efficient mining demands, and associated hazards of using explosives, it is necessary to utilize explosives at their best. Efficiency of a blasting operation is generally evaluated by the fragmentation size. From post blast conditions at blasting site to efficiency of processing plant, anything can be affected due to improper blast resulting in bad fragmentation. Better fragmentation is necessity for better operation of the system along with consideration of the safety of operation, environmental issues and related costs. For developing a blast design, an engineer should always attempt to minimize the cost of operation without compromising with acceptable technical and safety necessities [1]. Blast performance is dependent on many different factors associated with the process. These factors have been gauged as blast geometry, explosive specification, and rock mass properties [2]. Similarly, in 2013 Mohammad et al. concluded that fragmentation dynamics depends on blast hole diameter, primer, spacing, stemming, burden, and delay timing for blasting operation [1]. At the work site where this study was performed, blast holes were filled by drill cuttings and muds to help in confining gases. In a blast hole, gas confinement helps in better rock fragmentation along with reduced flyrocks. This practice is most common in many countries including Pakistan. Due to lack of homogenized particle size, the effectiveness of confinement is variable and majorly based on the size of stemming. Extremely small sized drill cuttings often result in hole blow out along with flyrocks from the collar. This improper blast energy will result in poor blast, bad impacts on environment, and reduced operational efficiency. In addition to the stemming material, the length of stemming is also a prime consideration. The decrease in stemming length will increase the quantity of explosive in the blast hole and an increase in stemming material generally improves stemming performance.

Jan 1, 2019

By H. Dr. Boldt, J. Dr. Henneke

PREFACE In German coal mines the application of tunnel boring machines has become a very effective and economic technique for major development works. After initial trials around 1970 the breakthrough was reached by driving 13 km of road ways on Rheinland and Friedrich Heinrich collieries, using a DEMAG tunnel boring machine of 6 m diameter. INTRODUCTION The coal mines of Rheinland and Friedrich Heinrich are to be found in the West of the Ruhr Area, the largest coal mining area in the Federal Republic of Germany (Table 1). They form part of the Ruhrkohle company. Rheinland coal mines produces 41000 to 43000 t of run-of-mine coal every day, of which 20500 t are usable coal. Friedrich Heinrich has a daily production of 17000 t of run-of-mine coal, 9000 t of which are usable. The coal is worked at depths of 600 to 1100 m in seams 0.6 to 3.0 m thick. The method of mining is longwall, and the machines used are ploughs and shearer loaders. From 1973 to 1980 a programme was put into operation in these two mines to develop and open up the coal field on the 885 m level. A total tunnel length of 20 km was to be driven. After thorough investigations it was decided to employ a full-face tunnel boring machine for

Jan 1, 1981

By Sai Phaninder Reddy Chaganti, Corrina D. Beaven

This study investigates the effect of reagent type, dosage, and pH on flotation kinetics and copper recovery for secondary sulfide ore with elevated clay content. The primary objective was to determine whether Xanthate performs comparably to or better than the currently used primary collector, Thiophosphate. Approximately 80 kg of ore was collected from three blast holes and systematically prepared through drying, crushing, grinding, and flotation testing. Grinding tests were conducted to establish optimal grinding time and lime consumption for pH control, achieving a target particle size distribution in 13 minutes and 30 seconds. Elemental and mineralogical analyses using X-ray Fluorescence (XRF) and X-ray Diffraction (XRD) identified a copper content of 0.36% and significant clay presence, affecting flotation performance. Fifteen flotation tests were performed under controlled conditions, with variations in reagent dosages. Control tests confirmed reproducibility, with recoveries around 80–82%. Increasing the pH did not improve kinetics or recovery, making higher pH adjustments economically unjustifiable. Adjusting the Thiophosphate dosage yielded the best recovery of 83%. Xanthate collectors, tested both as tertiary additives and as replacements for Thiophosphate, delivered comparable or superior recoveries—up to 91.5% in plant tests—and demonstrated faster flotation kinetics, even under high-clay conditions. These findings highlight Xanthate’s reduced sensitivity to clay, potential for lower dosage requirements, and operational benefits such as improved selectivity and circuit efficiency, validating its role as a cost-effective and alternative for processing complex sulfide ores.

Feb 22, 2026

By Jacques H. Nantel, James F. Archibald

INTRODUCTION The use of backfill materials in underground mines has been practiced for many years and the trend indicates an increased utilization of backfill materials in the future. The authors believe that underground uranium mines will follow this trend. Classified mill tailings, approximately 65 to 70 percent solids by weight, are the most common fill materials. Cement or other pozzolanic agents can be added to the fill materials to improve their strength characteristics. Dry fill, in the form of development muck, can also be introduced in the stopes to fill the voids left by the action of mining the ore. This paper deals mainly with the assessment of the radiation properties of backfill materials originating from uranium ores processed in the milling facilities of uranium mines. The paper outlines the results of studies conducted in the laboratory and in the field on radon emanating characteristics of backfill materials utilized in uranium mines. ROLE OF BACKFILL IN UNDERGROUND MINES Fill is a structural component of any underground mining system. In general terms, the purpose of filling is to provide ground support, to make recovery of pillars possible to provide a working platform for men and equipment, and to allow for waste disposal. There are other inherent advantages to the use of backfill underground, the most important being to: a) permit mining at greater depth b) allow the adoption of highly mechanized mining techniques c) reduce dilution from wall sloughing d) minimize roof and wall failures e) allow the mining of irregular orebodies, and f) to provide better control of ventilation. The last item is of great importance to uranium mine operators. The introduction of radioactive backfill underground represents an additional source of radioactive contamination; on the other hand the introduction of backfill in the mine reduces the volumes of workings to be ventilated. As a consequence the general mine ventilation is more effective and in most cases, better air quality can be maintained in the mine. There are of course some disadvantages associated with the use of hydraulic fills in underground mines, notably: a) the handling of large additional quantities of water from the underground workings b) delays caused by the cyclical nature of the backfilling operations c) additional costs associated with stope preparation and placement of backfill in stopes, and d) "spills" caused by bursting pipes or fill escaping from fill retaining structures. ASSESSMENT OF NON-RADIOACTIVE PROPERTIES Although the radioactive properties of mine backfills are discussed in this paper many other physical and chemical properties of backfills must also be considered. These properties include: a) percolation rate and permeability parameters b) viscosity versus pulp density relationships c) size analyses d) settling times e) settled bulk density f) solids specific gravity g) void ratio and porosity h) optimal percentage solids by weight i) specific gravity of slurry j) addition of cement:cement retention and loss k) unconfined compressive strength 1) triaxial compressive strength m) shear strength (non-cemented)

Jan 1, 1981

By S. L. Manske

The Mesquite gold deposit is a recent discovery of Gold Fields Mining Corporation lying 65 kilo- meters northeast of El Centro in Imperial County, California (Figure 1). The deposit is located on the southwestern flank of the Chocolate Mountains and is a bulk-minable, low grade "disseminated" orebody. Although the host rocks of the deposit are part of a regional high-grade metamorphic terrane, abundant brittle fractures, brecciation, and open-space vein fillings indicate that Mesquite mineralization is epithermal. Faulting which postdates the hydrothermal activity has complicated the original geometry of the orebody. Alluvium effectively hides most surface expression of the mineralization, but open-pit mining is now revealing the details of Mesquite geology.

Jan 1, 1987

By Jerry Solomon

The subjects to be examined in my portion of the session entitled "Technology of Aluminum.. Recovery" will be twofold. 1. Economics of the basic aluminum economy--supply-demand-price. 2. Specifications of aluminum products produced from municipal waste--from the secondary aluminum standpoint. It is the object of this speech to first impress everyone with the excellent long-run markets for the aluminum products from municipal waste and then to acquaint everyone with the rather liberal scrap specifications that the recycling industry can tolerate. The most important statement that can be made about the recycling of aluminum is that it is now and will be an economic necessity. This is true far many reasons; one which hit the headlines recently was Interior Secretary Rogers Morton's statement "Third-world nations could use mineral wealth not to achieve political ends but to jack up economic positions. The result could be sky-rocketing prices and dwindling supplies on world markets." There is no reason why the group of countries that supplies most of our bauxite can't get together the way oil-producing countries got together on the price of oil. (Jamaica and Surinam, are original sources of two-thirds of the aluminum used in the United States.) The worry is not so much that we will end up without the basic raw material but rather that the price might be so high that we could not afford it. So much for the blackmail reason for recycling aluminum; the supply and demand picture in both the primary and recycling sectors of the aluminum economy is far more meaningful.

Jan 1, 1974

By Andrew Torok

Activated bleaching clays, so called because of their ability to remove color bodies from bleach fats and oils, are produced from clays of bentonitic origin. Activation is achieved by subjecting these clays to physical or chemical alteration to increase selective sorption capacities (1) . The future of activated bleaching clays depends upon the ability of producers to recognize the new areas of applications which are arising, and their skill in developing sophisticated active clays to meet new and stringent requirements. The major portion of activated bentonites are produced for decolorization of oils. New techniques of activation will improve these clays for broader spectrum use. General properties expected to result from present activation methods are: 1. Modification of hydrophobic-hydrophilic properties and increased available surface area in non-aqueous systems, thereby improving sorption capacity and ability to remove color bodies from oils. The following additional properties will be expected from future activated clays: 2. High purity of the active clay with interfering impurities totally removed. 3. Increased selective sorption capacity in aqueous systems. 4. Selective chemical reactivity, such as oxidation or hydrolysis. This property is required for removing

Jan 1, 1971

By Elena Chiriotti, Federico Valdemarin, Jan Cenek

"SUMMARY The “28 May Station” is the only transfer station between the existing Green and Red metro lines in Baku. As part of the huge development strategy of the metro network, the complete separation of the existing underground lines has been planned at this location, in order to simplify the operation of the future metro system. The extreme challenge of the project is represented by the need of achieve the separation of such lines over a breakdown of the operation of just five weeks. The lack of direct access from the surface, a complex geology with artesian water layers, a congested underground space and sensible buildings represent additional constraints. The chosen design solution consists in creating a complex steel pipe jacking frame around the operating tunnels and then excavating under such frame two large caverns where the separation of the existing lines and the construction of the new metro tunnels will be achieved. Soil treatments have been conceived to mitigate the geotechnical risks and to reduce possible impacts on the existing structures. The paper illustrate the challenges of this extraordinary project, one of the most complex underground structures under study, the design solutions and the risk management approach used to manage the complexity and to design the mitigations measures. INTRODUCTION The metro network in the capital of Azerbaijan is rapidly growing together with its metropolitan area. The beginnings of the existing metro system started in Baku in 1967 in Soviet times. To date, only two lines (Green and Red) are in operation with a total length of 34 km, including 23 stations. The future network will be much wider: besides the construction of three completely new lines and the extension of two existing lines, the whole metro project includes the development and modernization of the existing metro sections. A consortium formed by Systra (leader), Mott MacDonald (Praha) and Saman (Korea) awarded the contract for the preliminary and detailed design of infrastructures for the extension of Baku Metro Network. This extension consists in 83 km of new lines and extension of existing lines. As part of this modernization project, the connection of the Green Line at the 28 May / Cefer Cabbarli Stations, focus of this paper, represent also it’s ""Achilles' heel"" given that the already very congested system have only one transfer station."

Jan 1, 2016

By Howard J. Haselhuhn

The effectiveness of selective flocculation and dispersion processes for the concentration of hematite ore are strongly dependent on the ionic content of the process water. It has been noticed that magnesium ions are more detrimental to iron recoveries during gravity separation and flotation processes than calcium. This phenomenon was studied by measuring the zeta potential of hematite ore at various concentrations of sodium, magnesium, calcium, strontium and barium at a pH of 11. Results show that zeta potential inverts from negative to positive at concentrations as low as 0.16 mM of magnesium ions in the solution. It takes a significantly higher concentration (1.25 mM) of calcium ions to achieve the same effect. This difference is attributed to the ability of magnesium ions to adsorb to all surface hydroxyl groups, whereas calcium, due to its larger size, can only adsorb to every third hydroxyl group. This hypothesis was confirmed by results similar to those seen with calcium when this test was repeated with strontium and barium ions. The tendency for calcium ions to adsorb to every third surface hydroxyl group causes higher concentrations of calcium to be less detrimental to hematite concentration processes than magnesium.

May 1, 2012

By E. Oraby, S. Garcia, A. Thurgood, D. Bertoldi, G. O’Malley

Mining and Process Solutions (MPS) most advanced process, GlyCatTM is the only process that can be easily adapted to existing cyanidation plants in the market today whose real advantages over cyanidation are when treating copper containing ores, concentrates or tailings. With such complex material, GlyCatTM requires 70-90% less cyanide addition to achieve the same precious metal extraction. However, GlyCatTM is not simply cyanidation with added glycine. It is a completely different process where the solution chemistry and conditions needed to effectively leach the precious metals are quite different. Downstream however, the process is largely unchanged as the recovery of the leached gold and silver uses existing known processes such as conventional activated carbon or zinc cementation (Merrill-Crowe process). Another advantage is lower SCN, OCN, WAD and no free cyanide in the effluent which can substantially reduce cyanide detoxification requirements and decrease the OHS and ESG risk of the operation. This paper outline some of these differences and presents testwork, piloting and process examples to highlight the advantageous of this new process over conventional cyanidation.

Feb 1, 2023

By Edmund H. Warner

Any city dweller who has walked alone along a remote mine passageway has to be impressed by the eerie silence - the total absence of noise except for ones own breathing and scuffing of boots on the bottom. An underground mine can have tremendous contrasts, from complete silence to the ear piercing clatter of percussion drills working on the mine roof. Noise in underground mining machines is both useful and detrimental. Without characteristic noises that are recognizable to miners, operators would have a more difficult job operating their machines. Noise serves as a warning of approaching machinery, and changes in noise give indication that breakdowns may occur. But excessive noise is a health hazard. In recognition of this fact, our Congress included noise exposure as a part of the "Coal Mine Health and Safety Act of 1969". As Fig. 1 indicates, allowable noise exposure is established by specifying a time duration permitted at different noise levels. This is the well known "Walsh-Healey" criteria which governs noise exposure in many industries.

Jan 1, 1977

By Bernard L. Cohen

INTRODUCTION The most serious occupational radiation episode in U.S. history was the exposure of uranium miners to radon daughters in the 1945-1968 time period. Although the maximum permissible dose at that time was 12 WLM (working level months) per year, average exposures were several times that figure, and ten times higher exposures were not uncommon. Starting in the early 1960s, mounting epidemiological evidence appeared for an excess incidence of lung cancer among these miners; up to 1974, there was an excess of 134 cases (159 observed vs 25 expected) among the 4000 in the group under study (NAS-1980). When this excess became apparent, the Federal Government took jurisdiction, greatly tightened enforcement, and in 1969 lowered the maximum permissible exposure to 4 WLM per year. These standards were met largely by a great improvement in ventilation. As a result of these measures, average miner exposures were reduced 5-fold from 1965 to 1968 and by another factor of 3 by 1970. In 1978, the 4 WLM/year maximum permissible exposure was exceeded by only 37 of 7500 underground miners, for only 16 wars it exceeded by more than 25%, and for only 8 was it exceeded by 502 (AIF-1980). The average exposures to the various groups are listed in Table 1. We see that the average exposure for all miners was 1.03 WLM, and for those who worked essentially full time underground it was 1.45 WLM. For purpose of later discussion, it will be convenient to choose a single value for average exposures. In view of the fact that employment situations and job categories are bound to vary over a lifetime of work, we take this to be 1.3 WLM/year. The purpose of this paper is not to dwell on history, but rather to address the question of whether or not this present situation is satisfactory. This question was considered recently by a study group under the auspices of National Institue of Occupational Safety and Health (NIOSH-1980) and it answered the question in the negative, concluding that the maximum permissible exposure should be substantially reduced. However, their arguments were incomplete and were lacking in perspective. The principal thrust of this paper is to provide some of the material overlooked. RISK TO URANIUM MINERS FROM RADON EXPOSURE In order to quantify the risk to uranium miners under present working conditions, it is first necessary to estimate the risk of lung cancer per WLM of exposure. The most straightforward way of doing this is to use the data on the group of uranium miners under study. These are listed in Table 2 for 8 exposure ranges and for the total group (NAS-1980). If all data are given equal weight, the risk is seen to be 3.5 x 10-6 per year per WLM. Much can be said in favor of using this value for the risk as dose independent; it is within one standard deviation (SD) of the observations in 5 of the 8 dose categories (as expected from the definition of SD), and within 2 SD for all 8, and for the three cases differing from the mean by more than one SD, two are above and one is below. However, the uranium miner exposures in the present situation will be far below those covered in Table 2, so it is worth considering the possibility that the risk differs from this at low doses. From the last column of Table 2 we see that the experience from all exposures below 600 WLM indicates a risk higher than the mean for the entire group by more than two standard deviations (the value 600 WLM in this comparison was deliberately chosen to maximize this deviation; for exposures below 360 WLM and 840 WLM, the excess is only 1.0 and 1.6 standard deviations respectively). In view of this tendency for the risk to be higher at lower dose, we will take the risk to be 5.0 x 10-6 per year per WLM in what we will refer to as model A. This risk estimate is based entirely on exposures much [higher] than those that will be experienced by

Jan 1, 1981

By F. P. Schwarz, A. J. Erickson, S. M. Weber

The Mount Hope stockwork molybdenum exposit, Eureka County, Nevada, occurs in a small igneous cauldron complex 32 to 38 mybp in age. The deposit is characterized by alteration zones and mineralized shells symmetrically distributed around multiple intrusives. For samples from the 25 000 meters of 1981-82 drilling, an appraisal of error in in-house sample preparation and in commercial assaying has been made. After splitting, drill core is crushed and reduced in volume (8:1 at 10-mesh, 8:1 at 70-mesh, 6:1 at 200-mesh) to supply 25 grams of 200-mesh pulp for replicate analyses. A 0.5 g portion of the pulp is used for dissolution and atomic absorption analyses. Factors which initially generated unacceptable data variances have been minimized by planning, repeat analyses, introduction of standards, and routine monitoring. Assuming that error between replicate assays defines a Gaussian distribution with its average ideally equal to zero, precision (from preparation including core splitting and analysis of 8400 pairs of molybdenum samples) is on the order of + 25%. The error distribution is not biased. Accuracy is approximately 4% lower than absolute as determined by replicate assays of commercially prepared samples.

Jan 1, 1992

By Douglas B. Silver

Study Parameters (Status as of February 1993) Time Period: 1989 -1991 151 North American Public Metal Mining Companies Fiscal not calendar year basis 85+% of Canadian and U.S. Metal Production Au, Ag, Cu, Pb, Zn, Mo, Ni, W Results are Preliminary and Research Continues Chicken Little is Alive and Well

Jan 1, 1993

By Barry K. Johnston, Norman W. Bainbridge, Neville C. Lockhart

CSIRO and its collaborators have developed a major R & D project on dewatering of coal. This involves sub-projects on (i) the fundamentals of coal-water associations; (ii) reducing the variability of product moisture levels from small coal centrifuges; (iii) process mechanisms and optimisation for fine coal dewatering; (iv) pilot scale testing, engineering development and innovation. Results from each of these sub-projects are presented, and the pilot facility incorporating a vacuum filter, belt press, membrane press, hyperbaric filters and centrifuges, is discussed. The vacuum filter and membrane press can be configured for electric-field assisted dewatering, which provides substantial enhancements in the rate and degree of dewatering for fine coal, coal tailings, and other suspensions.

Jan 1, 1995

By B. Qiu, Y. Luo

The design of the predriven recovery room is not an easy task, because the stress distribution in the barrier pillar, standing supports, fender pillar and the immediate roof is changing as the longwall face approaches the predriven recovery room. An analytical approach is proposed to determine the stress distribution in the barrier pillars, fender pillars, standing supports and immediate roof of the recovery room based on the fracture mechanics approach. Based on the developed mathematical model, a stand-alone computer program has been developed to facilitate the required computations. The parameters such as overburden depth, recovery room width, mining height, shield and standing support capacities are incorporated into this program to access the feasibilities of predriven recovery room designs. Compared to the finite element method previously employed, this analytical approach offers the same, if not better, accuracy without requiring expensive computer hardware and software or specialized expertise.

Jan 1, 2012

By Neall. George M.

No method of roof control yet devised has proven to be universally acceptable for the wide range of strata conditions experienced in U.S. coal mines. However, a relatively new innovation, the roof truss, has proven itself capable of preventing roof failures which both conventional and resin bolting could not. Presently, the economics of roof trusses are not favorable when compared to bolting due primarily to drilling and installation procedures, but this is expected to change with improved truss and equipment designs. To realize the full potential of this new method of roof control, design criteria are needed which can optimize the effect of the truss for varying strata conditions. A review of the literature reveals that some work has been conducted on the roof truss. 1 Gambrell and Haynes (1970) showed that the roof truss is superior to angled roofbolts installed through a header. Sheorey, Verma, and singh2 (1973) used statics to prove that the optimum angle of inclination should be greater than 450? However, to date, no systematic attempt has been made to establish design criteria. In this report, the problem was approached in three ways. First, model studies were performed to determine general trends of truss support as a function of a number of variables. This provided a better understanding of the support mechanism involved. Secondly, beam theory was employed to calculate the actual support capacity of a truss, realizing the many simplifications, assumptions, and shortcomings of this approach to such a complicated problem. Last, field testing will be utilized to test the hypotheses previously developed and provide hard data to further synthesize roof truss design criteria. Field testing has recently commenced and should be complete before mid 1977.

Jan 1, 1977