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By Henry T. Miller

INTRODUCTION The basic purpose of this paper is to suggest procedures and practices to insure that no employee working in a uranium milling operation receives exposure to radioactive, toxic, or other materials or agents that might produce a permanent, deleterious effect upon his physical health and well-being. This program is also designed to insure that each employee can carry out his assigned duties without risk to his health or to that of his fellow workers. The assumptions underlying this program are by necessity, conservative, with emphasis being placed upon prevention rather than correction or reaction to an undesirable situation. The limits specified-either as Threshold Limit Values (TLVs), Maximum Permissible Concentrations (MPCs), or Radiation Protection Guides (RPGs)--represent both standards of best practice and those of law and regulation. The total program is envisioned as a balanced combination of occupational hygiene and radiation monitoring. This includes surveying, air sampling, personnel dosimetry, bioassay, medical surveillance, epidemiology, and training--all backed by a thoroughly tested and evaluated set of emergency procedures. The program, as presented, is keyed to the results of monitoring, surveying, air sampling, medical surveillance, and epidemiology--it being obvious that no problem can result when no hazard can be identified. The lack of an identified hazard, however, does not negate the need for such a program and prudence suggests the need for routine surveillance programs in the best-designed facility. [The Model Mill] The model mill used in this paper is that presented by the U. S. Environmental Protection Agency in its study of the uranium fuel cycle1. The model mill processes 600,000 metric tons of ore per year into 1,140 metric tons of U308 as yellowcake per year using an acid leach process. Other characteristics of the mill such as location, tailings dam design, number of light water reactors supported, and geographical location are not pertinent to this paper. The usual remote western location of these facilities produce some difficulties in delivering the required occupational medical services and is not within the scope of this paper. The operating lifetime of most mills is commonly estimated as from 12 to 15 years in actual practice; however, the U. S. Environmental Protection Agency estimated a 20-year lifetime for its model. The principal steps in the acid leach process include :2 (a) Ore is blended and crushed to pass through a 2.5 cm (one inch) screen. The crushed ore is then wet ground in a rod or ball mill and is transferred as a slurry to leaching tanks. (b) The ore is contacted with sulfuric acid solution and an oxidizing reagent to leach uranium from the ore. The product liquor is pumped to the solvent-extraction circuit while the washed residues (tailings) are sent to the tailings pond or pile. (c) Solvent extraction is used to purify and concentrate the uranium. (d) The uranium is precipitated with ammonia and transferred as a slurry. (e) Thickening and centrifuging are used to separate the uranium concentrate from residual liquids. (f) The concentrate is calcined and pulverized. (g) The concentrate or yellowcake is packaged in 208 liter drums for shipment. This process is illustrated in Figure 1.3

Jan 1, 1981

By Jerzy Iskra, Wlodzimierz Sikora

INTRODUCTION The effect of stress concentration in coal seams leads to a sudden release in the form of rock bursts. This effect is controlled along many lines, but the most efficient methods consist in preventing the accumulation of energy in the seam or in the roof. The over- burden weight resulting from mining is the main cause of excessive stresses in a seam. An additional weighting of roof takes place due to formation of abutment loads. The accumulation of energy in the roof takes place where thick sandstone beds occur in overburden strata. An essential problem is to watch, with appropriate measurements, the stress of coal and roof rocks, and to conclude, on this basis, the degree of rock burst hazard. The methods for determining the stresses in a coal seam are quite well developed at present. Let me mention here the seismoacoustic methods, or borehole overcoring measurement methods. On the other hand, an examination of the state of stress in the roof rocks is almost impossible at the actual stage of development of the measuring technique. Hence, the model of stresses in roof was analyzed and determined theoretically. Based on this model, a technology for roof fracturing was developed which aims at destressing rock masses in a planned way. The essentials of the fracturing method consist of a controlled breaking of sandstone roof beds in order to: - prevent the excessive weight on the pillar - prevent a spontaneous breaking of roof beds at a limiting state followed by sudden energy release. High pressure fluid infusion is used to release buildup of stresses. THEORETICAL AND RESEARCH BASES According to the tests carried out by A. Kidybinski by the method of finite elements, isobars of stresses occurring in the roof over the working face are shown in [Fig. 1]. The results indicate: - the highest values of compressive stresses occur in the roof directly over the seam, at a distance of about 5 m ahead of the working face - at a distance of about 22 m above the seam there is a line of transition of compressive stresses into tensile - within the tested roof area, the highest tensile stresses occur at a vertical distance of 35 m, and at a horizontal distance of 20 m from the line of the working face. H. Gil (1) determined the pres- sure initiating the fracture of roof rocks by the formula: [ ] On the basis of results of analysis of isobars, and of the initiation of failure stress formula (1), isobars may be designed for a given roof class. Assuming for sandstones the Poisson ratio equal 0.3, a field of failure stresses presented in [Fig. 2] is obtained. In this figure it can be seen in which areas the effects of sandstone breakage may be obtained with the operation of a determined hydraulic pressure. To use formula (I), the knowledge of Rr is still necessary. This value may be determined in two ways: - taking a sample and testing it in the laboratory, if there is access to roof rocks in the existing openings, - drilling boreholes in the roof from surface and examination of the rock mass by using hydraulic breaking, if there is no access to roof rocks.

Jan 1, 1982

By R. K. Mehta, S. K. Dhar, R. R. Mallepali

Introduction The incentives for developing a distributed form of the comminution model include: • for describing the heterogeneous comminution behavior of multiphase mineral systems (i.e. realistic ore bodies, such as complex sulfide ores consisting of chalcopyrite, galena, sphalerite, pyrite and siliceous minerals) and •for generating, for various locked and free particles, in situ, microscopic, breakage behavior such as the mode of fracture (intergranular vs. transgranular), the relative rates of breakage and the breakage distribution. However, the computational difficulty and the effort made in estimating a large number of parameters (Andrews and Mika, 1975; Wiegel, 1976; Herbst et al., 1985, 1988) do not justify the utility of this approach. Also, obtaining the results experimentally is very tedious and time consuming if using this approach. Although recent work in this area (Choi, Adel and Yoon, 1987; Mehta, Adel and Yoon, 1989, 1990) showed good results for liberated particles, the quality of the estimation and prediction was poor for locked particles. Attributing this to the complex breakage behavior of locked particles, it was concluded that additional constraints, and possibly modification in the breakage distribution functional form for locked particles, were necessary. However, further mathematical analysis revealed that the method previously used was in error. The purpose of this technical note is to present the correct solution methodology used in solving the general model equations of a multicomponent mineral system containing multiple classes of locked particles. Previous work and connection The general form of the comprehensive, distributed-component comminution model has been described as (Herbst et al., 1985) [d (m (0) -[ I - B ] S_ m(t) = A m(t) (1)] where [A] is the fundamental matrix/state companion matrix, commonly referred to, in the comminution literature, as the "mill matrix." Note that for a distributed-component comminution model, this is a full, general, real matrix, unlike a lower triangular matrix for a simple grinding model or a lumped-binary grinding model with one locked-particle class (Herbst et al., 1985, 1988). Also note that all of the underlined letters refer to matrices. Herbst et al. addressed the difficulty in arriving at a closed form of the analytical solution of Eq. (1) via similarity transformation [IT J inv (T)]]. This difficulty occurs for several reasons: •[A], as a general real matrix, may have complex eigenvalues and complex eigenvectors. •More importantly, there is no guarantee that a set of eigenvectors of [A] will turn out to be independent. An attempt was made to solve this problem by implementing a decomposition method. With this technique, the mill matrix was decomposed into a sum of lower and upper triangular matrices [(i.e., A = L + U)]. Then the solution was obtained using similarity transformation (Mehta, Adel and Yoon, 1989, 1990). [e(At) = e(L+U)t = e(Lt).e(Ut) _ [ TL Jt- (t) inv(T_L)] [-Tu Ju (t) inv(Tu)]] [But, unfortunately, e 1+U = e t . eu is true only for scalars and not for matrices (i.e., eL+u# eL- eU) unless they commute. In other words. eAt = e(L+u)t= ekt - eU' if LU = UL. Hence, this decomposition scheme must be replaced by a mathematically acceptable method known as the "splitting method"]

Jan 1, 1994

By R. K. Zipf

Recent scientific studies by researchers at the National Institute for Occupational Safety and Health (NIOSH) and the U.S. Army Corps of Engineers (USACE) presented the worst-case explosion pressures that can develop if a methane-air mixture of a certain composition and dimension accumulates and ignites within a sealed area of a mine. The Mine Safety and Health Administration (MSHA) issued its final rule on sealing of abandoned areas which considers potential explosion pressures and the risk of achieving the necessary conditions. The new MSHA seals regulation has increased the pressure design criteria an order of magnitude beyond the old standard of 140 kPa (20 psi) to 800 kPa (120 psi). However, there still are major knowledge gaps pertaining to seals and sealed areas including: (1) the composition of the sealed area atmosphere and how it changes in time and space, (2) explosion pressures that might actually develop within sealed areas, (3) engineering procedures to follow when designing seals and seal installations, and (4) guidelines to manage the atmosphere within sealed areas with monitoring or inertization. NIOSH has initiated a seals research program with collaborators from Safety in Mines Testing and Research Station (SIMTARS) in Australia, West Virginia University (WVU), the Naval Research Laboratory (NRL), and USACE. Highlights of progress so far include: (1) construction of a gas explosion tube at Lake Lynn Laboratory for studying large methane-air explosions, (2) computer simulation of the fluid dynamics of high pressure methane-air explosions, (3) development of seal analysis methods considering construction materials, reinforcement and the seal foundation, and (4) development and testing of innovative seal designs that utilize inexpensive, readily-available rock rubble to resist high explosion pressures. NIOSH researchers aim to provide sound engineering guidelines to better address the new MSHA seal regulations. DISCLAIMER: The findings and conclusions in this report have not been formally disseminated by the National Institute for Occupational Safety and Health and should not be construed to represent any agency determination or policy.

Jan 1, 2009

By Nicola Beltrandi, Andrea Pigorini, Holad Mohamed Dahir, Ezio Pedone

"Genoa is a very distinctive city located between mountains and the sea and infrastructural upgrading works often have to deal with the intersection of existing infrastructures. This is the case for the new railway tunnel construction between the stations of Brignole and Principe which under-cross two existing double-track tunnels in operation. The latter have been constructed with masonry bricks and the vertical distance between old and new tunnels is approximately 1m. Due to rock characteristics and the urban environment, stability and vibrations control are the main issues. Consolidation work with the fore-poling method and a specific sequence of excavation, primary lining and final lining support have been designed to assure stability and settlement control. A hard rock road-header has been used to assure vibration control. Finally, a real time monitoring has been implemented. The article presents monitoring and back-analysis results of the under-crossing of the existing tunnels. GENERAL LAYOUT OF WORKS The Genoa railway junction constitutes an integral part of the Ventimiglia-Genoa-Novara-Milan railway line route and it is included amongst the railway systems of the Tyrrhenian-Northern European multi-modal corridor (corridor of the two seas or corridor VII Genoa-Rotterdam). The main aims of the project concern the upgrading of the existing lines and the separation of the long-distance passenger and freight traffic flows from those of a regional and metropolitan type. The project includes: •,Completion of the connection with the port of Genoa Voltri and the Italy-France coastal route. •,Interconnection with the future Terzo Valico dei Giovi mountain pass route towards Milan. •,Upgrading of the Genoa Voltri – Genoa Brignole main line to four/six tracks. Rearrangement of the systems at the Genoa Voltri, Genoa Sampierdarena and Genoa Brignole Stations, reorganisation of the services in the Genoa Terralba area and interchangeability with the metropolitan system complete the design activities. The contract which includes civil and technology works is divided into four construction lots (ref. Figure 1):"

Jan 1, 2016

By G. W. Cutting

It is well known in both the research field and in commercial flotation processes that the grades and recoveries obtained from flotation plants are heavily influenced by the structure and height of the froths produced in the flotation cells. The development of steady-state mathematical models of flotation processes based upon practical and readily obtained parameters have demanded detailed information about the structure of flotation froths. The important parameters are the distribution of floating values, gangue and water in the froth with respect to froth height, cell pulp density, cell air flow rates and flotation inducing additives. Little work has been conducted in this field although considerable attention has been paid to bubble loading and other factors close to the cell froth/pulp interface. Laboratory batch tests are prefaced for the derivation of the above mentioned parameters for both practical and economic reasons. The standard laboratory flotation tests using a batch cell produces a froth of low height (1 to 3 cms generally) together with a varying cell pulp density as the test progresses due to the addition of top up water and the removal of froth from the cell. This form of test can be considered only as giving information about froth conditions close to the cell froth/pulp interface under conditions of widely varying cell pulp densities. In commercial processes the froth heights are much larger (frequently greater than 10 cms); therefore the laboratory batch test, although useful in many ways, does not directly yield parameters relating the state of the froth flowing over the cell lip to process operating conditions. To investigate flotation conditions, where more realistic froth heights are developed, an equilibrium type of batch cell has been developed at Warren Spring 1 Laboratory . In this a tank with high walls is used to contain the pulp and whole froth and, after a short period of operation, an equilibrium between froth

Jan 1, 1975

By Ian P. G. Hutchison

12.1 INTRODUCTION This chapter deals with the tools required to design leach operations to minimize spillage from leach pads and ponds and erosion damage during flood events or prolonged wet periods. It outlines methods used to design for prolonged dry periods when water shortage can be a problem. Seasonal operational aspects such as winter shutdown and coping with spring snowmelt periods are also discussed. General criteria useful when designing pregnant and barren solution ponds are also outlined. The aim of these notes is to provide designers and operators with a simp1ified approach to developing design parameters which make effective surface water management easy to achieve without resorting to extensive or elaborate analyses. Section 12.2 briefly outlines the types of problems associated with wet climates and floods, as well as dry periods. Determination of flood design parameters is discussed in Section 12.3. Circuit water balance and pond sizing are outlined in Section 12.4. Pond design considerations are included in Section 12.5. Worked examples are included in Section 12.6. 12.2 RAINFALL/RUNOFF PROCESSES There are two types of rainfall-runoff processes that are important when designing heap leach facilities. The first of- these is the short-term, intense rainfall or snowmelt events that occur over a period of a few minutes or hours, causing flooding that can lead to erosion damage and overtopping of ponds and containment dikes. Specific problems that can manifest themselves during these events include: • Washout of access roads; • Damage to containment dikes; • Flood water contamination of circuit water; • Damage to process water ponds by perimeter erosion or overtopping and erosion; • Loss of process circuit fluid due to breaks in containment dikes or process water ponds; and • Erosion of leach pile and consequent sediment contamination of process fluid. Other types of rainfall -runoff processes that could cause problems are the long-term-(generally months-long) wet or dry periods. Extended wet periods

Jan 1, 1988

By G. W. Kenzy

Recently, in the United States, questions of possible deleterious synergistic health effects of the gaseous and particulate components in the diesel exhaust and mineral and coal dust have been the impetus for extensive research by several agencies. At the present time, diesel engines are tested and approved for use in gassy and non-gassy underground metal and non-metal mines under Parts 31, 32, and 36 of Title30 of the Code of Federal Regulations. Diesel-powered equipment for use in underground coal mines, though not explicitly addressed in the law, is administered under Parts 32 and 36 of Title 30, with Part 36 being applied in the case of equipment working inby the last open crosscut or in the return airways. Operation of diesel equipment in underground mines has the potential for serious health and safety hazards from the gases, vapors, and particulates in the exhaust. Therefore, various countries have rules and regulations for the safe deployment of diesel equipment. The applicable laws and regulations from several countries who have long experience with diesel-powered equipment in their mines are reviewed. Four major areas have been addressed: (1) testing and approval procedures of diesel engines, (2) ambient mine air quality specifications, (3) frequency of sampling and location of samples, and (4) air quality specifications in splits where diesel-powered equipment are operated. The review indicates that most regulations generally consider four possible hazards: (1) toxic components in the engine exhaust, (2) ignition of flammable atmospheres by the engine or electrical equipment, (3) fire hazards presented by the engine fuel oil and by combustible material that might come in contact with the equipment, and (4) mechanical hazards. There is a remarkable similarity in the laws with regard to engine design, testing and approval procedures. However, the aspects with regard to mine air quality and quantity are quite variable-- from a general statement to very specific standards for many different diesel exhaust constituents. However, the laws with regard to synergistic effects of diesel exhaust components and other mine air contaminants are an exception.

Jan 1, 1980

By S. S. Sephton

The La Perla Mine, owned by Altos Hornos de Mexico, S.A., is located approximately 100 miles east of the City of Chihuahua in the State of Chihuahua, Mexico. The mine began operations in 1958. Existing ore beneficiating facilities, consisting of heavy media and jig plants, upgrade the ore prior to shipment by rail to the Altos Hornos steelmaking complex in Monclova. Sizable reserves of two distinct ore types, which were not amenable to treatment in the existing beneficiating plant, or usable in the blast furnaces, have been mined and stockpiled in order to uncover the treatable ores. One of these ore types has been labeled "Diseminado" ore, and is low grade containing approximately 42% Fe with sulphurs ranging from 1 to 2%. The other ore type has been labeled "Pulverulento" ore, and is high grade containing approximately 65% Fe and 0.1% S, but at 80% minus 325 mesh, is too fine grained for efficient sintering. Late in 1968, Arthur G. McKee & Company became involved with the problem of improving the above ores for their efficient and economical use in Altos Hornos' blast furnaces. Early in 1969 representative samples of Diseminado ore were obtained for laboratory bench-scale testing at the University of Minnesota, Mineral Resources Research Center. Concentration tests incorporating magnetic roasting, wet magnetic separation, high intensity separation, high tension magnetic separation, anionic flotation of apatite, cationic flotation of silica and oleic acid flotation of iron were performed. Laboratory testing was concurrently carried on at La Perla Laboratory, which resulted in plans to construct a 10 TPD pilot plant. The pilot plant was operational in September, 1969. Analyzing laboratory results as well as pilot plant results, Altos Hornos and Arthur G. McKee & Company metallurgists agreed that the most promising flowsheet for the low grade Diseminado ores entailed wet magnetic cobbing, desliming, and oleic acid flotation of the non-magnetic fraction. A combined concentrate of

Jan 1, 1972

By Daniel Su

This paper describes recent ground control challenges encountered at the longwall recovery areas of one of Bailey Mine?s two longwall faces and the proactive measures taken by CONSOL management and engineering to mitigate them. The challenges derived primarily from the following causes: (1) irregular roof coal and draw slate thickness, (2) presence of thick sandstone at the recovery area and extremely weak rocks below the sandstone, (3) close proximity to a regional anticline, (4) low shield set pressure and tip load, and (5) ineffective primary roof support that led to progressive failure of the headgate roof. The presence of the anticline contributed to extensive fracturing of the sandstone. Excessive pot-out in advance of the shields frequently occurred near mid-face due to the presence of thick draw slate and irregular roof coal near the recovery area. This, in turn, resulted in face heights that exceeded the maximum reach of the shields. Subsequently, during longwall meshing, when the advance rate slowed considerably, fractured sandstone would often break loose, causing the shields to yield, thereby reducing the height and inhibiting shearer movement. To improve the headgate and face conditions within the longwall recovery area, several actions were taken, including: (1) assessing site-specific roof geology by conducting in-mine roof coring around the recovery areas and by mapping immediate draw slate and roof coal thicknesses within surrounding entries, (2) increasing pump pressure and thus shield set pressures (3) diligently monitoring shield set pressures and longwall extraction horizon, (4) installing additional cable bolts and roof screening to prevent progressive failure of the immediate headgate roof, (5) fully-grouting supplemental cable bolts to improve headgate support stiffness, and (6) gluing the headgate roof within the recovery areas with polyurethane. As a result of these collective efforts, longwall recovery delays at Bailey Mine were reduced from 21 days in 2006 to nearly zero in 2009.

Jan 1, 2009

By A. Gupta, S. C. Srivastava, K. K. Saroj

Bacterial degasification of methane and conventional methane drainage for generation of power have become accepted practice. Comparative economies of both the techniques have been described. Authors have discussed an alternative technique, involving complete methane biocycle. The first stage of the entire process involves gasification of coal using a consortia of bacteria to generate methane. The generated methane has been proposed to be pumped,as in conventional methane drainage. Technique described is basically similar to underground coal gasification, with the difference that in place of coal combustion, degradation of coal to lower molecular weight compounds by one set of bacteria and then methanogenesis by methanogens, producing methane have been described. Brief account of communities of bacteria that may be employed have been dealt with. Three arguments in favour of proposed technology as an alternative to conventional coal mining are: 1.Total elimination of hazardous operation. 2.Availability of power plants based on methane drainage, assuring comparative economies. 3.Possibility of utilisation of fuel cell technology for power generation giving higher efficiency. All the three phases of complete blocycle viz (1) methane generation from coal (ii) conversion of methane to hydrogen and oxygen (iii) conversion of products so obtained to methane, thereby completing the cycle have been discussed. INTRODUCTION Drainage of methane by conventional pumping process (Boxho,et.al.,1980) not only is helpful in maintaining safe extraction of coal, but with utilisation of extracted gas for power generation, also helps in over all economy of coal mining. This is gradually becoming standard practice for mining of coal in highly gassy mines. An alternative technique (Thakur et. al., 1991) of bacterial degasification of coal seams is also gradually gaining ground. The bye products of this process that is carbondioxide and water, do not contribute towards economy of coal extraction process. However the technique has its own advantage particularly in goaf area for avoiding fire hazard and drainage of methane from coal seams prior to initiation of mining operation. Either of these techniques could be applied in specific situations with success.

Jan 1, 1993

By Henryk Bystron

The underground fire areas are formed as a result of failure of an active fire extinguishing in mines. Such areas bring about a fire hazard for the work safety and for the mining operations. For liquidation of fire areas it is necessary to create conditions under which a possibly small amount of air is flowing to the fire. To achieve this end, isolation methods are used which consist in erection of full stoppings of various types, sealing practically the fire areas. These methods usually fail in oases when the coal seam is thick and the fire area is situated between the intake shaft and the outtake shaft in which a high-pressure fan is in operation. In such oases ventilation methods based on the balance of aerodynamic potentials in fire areas are to be used. For a very long time the isolation and ventilation methods have been developed in the world mining, No sooner than since 1980 the aerodynamic potential has been used in ventilation methods [1,2,6]. So far, the aerodynamic potential was called in different ways and was determined by different formulae which gave its numerical values more or less approximative. Nevertheless, even in situations featuring by a lack of sufficient theoretical bases of the aerodynamic potential the ventilation methods brought considerable profits in the field of safety in mines and of economics in mine ventilation [1-6,9,10]. The increase in the depth of underground mines and their equipment with expensive machines (apparatus) brought about that the Central Mining Institute at Katowice undertook research works which are crowned by the theory of the aerodynamic potential in mine ventilation networks [7]. The use of this theory in the mentioned ventilation methods will increase their effectiveness. Since the theory will be published in the III rd quarter of this year, we cite in the present paper the formulae and relations associated with this theory which are indispensable to present the balance of aerodynamic potentials in the fire areas in their control by ventilation methods.

Jan 1, 1980

By Omar A. Muhtadi

6.1 INTRODUCTION The three most critical factors in successful heap leaching operations are: • The methods employed to pretreat ore (i.e, crushing and agglomeration); • Placement of ore on the leach pad (heap building techniques); and • Wetting of the ore with cyanide (solution application techniques). Improper execution of any of these initial steps may result in a substantial sacrifice of total metal extraction or, in the worst case, in project failure. Ore preparation in terms of crushing and agglomeration was discussed in Chapter 5; this chapter therefore focuses on commonly employed techniques for heap construction and solution application. The main objective of heap construction is to obtain the most homogeneous heap possible. "Channeling" of leachate in heaps with preferential seepage paths leads to preferential leachate distribution and uneven/incomplete extraction. The two main contributing causes of channeling are inadequate attention to heap construction and uneven solution application. These pitfall s can best be avoided through a knowledge of the various techniques available and the applicability and success of these techniques with various ore types and site conditions. The remainder of this chapter provides a discussion on these two subjects. 6.2 HEAP BUILDING TECHNIQUES Successful heap building techniques vary with the type of ore and, since no two orebodies are exactly alike, some slight modifications are usually necessary for each specific project. However, it is possible to generalize in terms of the appropriate techniques to be used with ores having similar characteristics. This section discusses the three most commonly used methods of heap construction and provides some guidelines with regard to ore types, construction constraints, and lift height parameters for each of the three. 6.2.1 Run-of-Mine Dumping and Dozing The run-of-mine method of heap building can only be used with ore of a highly silicious nature (i.e., ore which will not generate a large amount of

Jan 1, 1988

By Gregg R. Bierei

Arch of Illinois is a mining complex comprised of two surface and two underground mines located in Perry County, in southern Illinois. The company a division of Apogee Coal Co. of St. Louis, MO. Arch of Illinois' two surface mines, the Captain and Horse Creek Mines, use large stripping shovels and cross-pit bucketwheel excavators to remove overburden from three seams of coal, the Herrin No. 6, Harrisburg No. 5 and Danville No. 7. All are considered high sulfur reserves. The two underground operations are managed by Arch of Illinois. They are the Kathleen Mine and the newly constructed Conant Mine, which began production in 1991. Both underground mines are continuous miner operations. They mine the No. 6 coal, which is about 1.8 m to 2.1 m (6 to 7 ft) thick. All coal is currently transported by truck to a central preparation plant located at the Captain Mine. This preparation plant produces 27 kt/d (30,000 stpd) of clean coal. Production in 1992 reached 5.9 Mt (6.5 million st) of saleable coal. Of that, about 66% or 4 Mt (4.4 million st) are produced by the surface mines and the rest the by under ground operations. Change is in the wind for Arch of Illinois. In 1988, 100% of all production (6.2 Mt or 6.9 million st) came from surface operations. By 1995, only 38% of production is scheduled from surface mines. And by the year 2000, 100% of production will from underground mining methods. Dwindling low ratio reserves, compounded by increasing environmental and land acquisition costs, coupled with more efficient underground mining equipment, have started an evolutionary trend toward development of underground reserves. For Arch of Illinois to maintain profitability of remaining surface reserves and to compete effectively mining underground, it must improve efficiency, productivity and cost reductions. These efforts are essential for maximizing current surface reserves and to ensure a continued existence in the presence of clean air legislation and the softening of the high sulfur coal market.

Jan 1, 1993

By Jennifer B. Leinart

With the signing of the North American Free Trade Agreement (NAFTA), Canada, the United States and Mexico entered into an international experiment in economics amid a cloud of controversy that still lingers. Central to the controversy are the disparities in wages that exist among workers in the signing countries. Wage surveys completed by Western Mine Engineering show the degree of disparity that exists among the mining industries of those three countries. Labor costs constitute 30% to 50% of operating costs. So these wage differences are a matter of interest to the mining industry. It is unclear how much of a factor they are in determining where companies search for and develop new mines. That depends on where minable deposits are expected to be found. However, a deposit that may not be minable in the United States or Canada may be minable in Mexico because of lower labor costs. Some adjustments are anticipated for the Mexican mining industry if and when its wages begin to approach comparability with US and Canadian wages. Table 1 combines the results of four recent surveys. It lists benchmark wages for mining job titles common to all the countries. About 220 mines are covered by the combined surveys. Canadian miners' wages, converted to US dollars, are typically 75% to 95% of US wages, while Mexican miners' wages are less than 10% of those in the United States. Table 1 also shows that wage disparities are not confined to international borders. In fact, the difference in average wages between the US coal industry and the US metal and industrial mineral industries is greater than between the Canadian and the US mining industries overall. Table 2 shows that managerial, technical and administrative salaries show similar, but less pronounced differences among the three countries. Salaried personnel at Canadian mines earn 80% to 90% as much as their US counterparts, while those in Mexico earn 20% to 30% as much. Here again, average US coal salaries exceed average salaries for the US metal and industrial mineral industries.

Jan 1, 1998

By William J. Douglas

Defining Productivity Productivity is a principal concern to all organizations - industrial, commercial, government, and academic. As a ratio of output to input, productivity measures the efficiency of an industrial process, much as thermodynamic efficiency measures the useful work obtained from the available energy consumed. However, there is one important difference. Whereas thermodynamic efficiency is a dimensionless a ratio of energies, productivity output and input are usually measured in different physical units. Nevertheless, the resulting ratio is a normalized expression for unit of output per unit of input to the process. In the early days of coal, mining was essentially a hand loading operation. There was little expenditure of alternative forms of energy within the process. Loading required expenditures of human energy on the order of 1.2 MJ/t (1,000 Btu per st). And if we accept the conventional value "you load 16 tons," the output to input ratio is 14.5 t (16 st) per manshift. In energy terms, it is on the order of 116 MJ/t (100,000 Btu per st) produced per labor Btu expended. However, the latter measure of efficiency was not meaningful in economic terms, while the 14.5 t (16 st) per manshift was significant. Since labor was the principal component of cost, it represented a direct relationship to the cost of production: [ ] Furthermore, when workers were paid directly on the basis of tons produced, the labor cost per ton was specified directly by this rate. With the introduction of mechanized loading, significant increases in productivity occurred, when measured on the basis of tons per manshift. Electric power consumed about 11.6-116 MJ/t (10,000-100,000 Btu per st). This makes the 29-GJ/t (25 million Btu per st) coal output/input ratio between 100 and 1,000. That is one or two orders of magnitude lower than hand loading. However, electric energy was quite inexpensive compared to labor, which made the mechanization economically efficient. The accepted productivity measure continued to be tons per manshift. This measure includes some of the effects of mechanization: namely, the additional manpower needed to maintain the machines and support the equipment that uses electrical energy. However, it does not consider the capital and operating costs that relate to mechanized equipment. As a result, the concept of tons per machine-shift has emerged as an accepted measure of shift productivity. The only appropriate measure that encorporates the productivity of men and machines is an economic one. The output is the tonnage per shift and the impact is the cost per shift. [ ] This is simply the inverse of the cost per ton. In addition, much flexibility is available in allo¬cating costs to the three categories. This adds a level of intangibility to the measure. The amount of information needed to obtain all of the required costs is appreciable compared to the simple matter of counting tons and workers. Furthermore, financial information relating to production costs is closely held by mining companies, so historical statistics on economic productivity are not publicly available. To extend the concept further, consider the value of the tonnage produced, to obtain a dimensionless ($/$) measure of economic productivity.

Jan 11, 1983

By J. F. Strunk Jr., K. M. Maestas, Y. Chai, R. Legrand, R. L. Henry

"Groundwater with elevated sulfate and metals was treated using a sulfate-reducing pilot PRB consisting of a series of six 55-gallon drums filled with ZVI (Fe0), wood chips, brewer’s grain, limestone, and emulsified vegetable oil (EVO). Sulfate was reduced from >3000 to <250 mg/L and Cd, Cu, Mn, Ni, and Zn concentrations declined up to 3 orders of magnitude in 2 months. After 6 months of operation at a 7.5-day retention time, sulfate broke through. The system was then rejuvenated by injecting EVO in each drum and operated for 5 more months at a 9-day retention time. Sulfate-reducing bacteria in the pilot system grew by 3 to 4 orders of magnitude in 2 months. Some cementation of the PRB media was observed in the first 2 drums after processing 25 pore volumes of water, due to preferential upstream precipitation of sulfides; this can be avoided through operational modifications. Resulting permeability losses were estimated at 0.1% to 1% per year. Absorption/desorption was tested at bench scale to evaluate the use of various oil-soaked sorbents as a slow-release source of dissolved carbon. The tests show that oil sorbents can increase the useful life of a PRB between oil refills to over 100 pore volumes. INTRODUCTION High sulfate concentrations, low pH, and dissolved metal exceedances are common characteristics of mining-influenced water (MIW). Above-ground such MIW is commonly treated using biochemical reactors (BCR). In a BCR, water generally flows vertically downward through a layer of biomass, such as wood chips, manure, etc. The dissolved organics resulting from the decay of this material are oxidized first by microbes using oxygen; when oxygen is exhausted, other bacteria can use nitrate, and then sulfate to oxidize organics. Sulfate-reducing bacteria (SRB) reduce sulfate to sulfide, which quickly reacts with metals to form largely insoluble and stable metal sulfides. These reactions can be summarized as follows:"

Jan 1, 2016

By C. I. Walker

"Understanding the influence of slurry particle size on wear is essential to maximising the life of centrifugal pumps operating in mineral processing plants and for hydrotransport duties. Particle size influences not only the wear rate but also the wear pattern in pumps. Complicating the situation is the different relative wear rate for different pump parts depending on particle size. A simple Slurry Wear Type (SWT) number based on particle size and size distribution is proposed as a means for categorising slurry pump wear patterns and optimising pump application and material selection. INTRODUCTION The wear of centrifugal slurry pumps in mineral processing and hydrotransport can be quite severe and is one of the major cost concerns of plant operators. Some typical worn slurry pump parts made from high chrome white iron are shown in Figure 1. Erosion wear rates of up to 2 mm/day are not uncommon [1].Centrifugal slurry pumps are required to handle solid-liquid mixtures with a very wide range of particles extending from sub-micron size clay through to 200mm and larger cobbles. Particle size distribution can also vary widely from uniform and narrow (such as graded sand) through to very broad (such as grinding circuit slurries). Slurry particles may also undergo size change due to attrition during the pumping process. The major wearing parts in a slurry pump are the impeller, side liners and casing liner. These are shown with reference to a Warman® MC centrifugal pump in Figure 2. While many particle properties (including size, shape, hardness, density and concentration) influence pump wear, this work will consider only the effect of particle size as it is the major determinant of wear characteristics. Particle size influences the wear pattern, material performance and relative component wear rate in centrifugal slurry pumps. SLURRY WEAR TYPE (SWT) BASED ON PARTICLE SIZE Soil and grain size classification systems abound. Many have different limits to the ranges for a particular classification, albeit not too"

Jan 1, 2018

By L. Giraldo, S. Cotten

"INTRODUCTION The vision for a “Mine Escape Vehicle” (MEV) is to provide capabilities to better enable underground travel to allow miners to escape to safety in the event of a major mine emergency such as a fire or explosion. Enhanced transportation and survival capability such as a MEV may have minimized or eliminated the loss of life associated with several of the mine disasters that occurred in 2006, which together resulted in the deaths of 19 miners. Although the most desirable response in an emergency is for the miners exit the mine, if escape proved to be impossible, the miners could use the vehicle to proceed to an underground point of refuge to await rescue. Under a NIOSH contract issued in April 2008, Raytheon developed a conceptual vehicle design including (1) the capacity to provide life support functions for a number of miners, (2) operate in a reduced oxygen atmosphere with little or no visibility, and (3) travel at speeds faster than miners could walk out of the mine. The proposed MEV concept envisioned modification of a personnel carrier (a.k.a. a “mantrip”) that would be used every day during normal mine operations, not a specialized vehicle reserved only for or dedicated solely to escape. The MEV concept also included escape aids that are detachable from the vehicle in the event that the vehicle must be abandoned on the way out of the mine and the dismounted miners had to proceed on foot. To advance MEV development, NIOSH awarded a second contract in 2010 to Raytheon to build, test, and evaluate the performance of a proof-of-concept MEV that culminated in a 7-month performance evaluation of vehicle in an Eastern Kentucky coal mine (Figure 1). User feedback and lessons learned from the proof-of-concept vehicle evaluation provided support for the general MEV approach, guidance related to the practicality of MEV capabilities, and indication of specific technology shortcomings for this application."

Jan 1, 2015

By J. H. Rowland, A. C. Smith, K. A. Teacoach

Current regulations for fire suppression systems in underground coal mines to protect conveyor belt installations have been in place since the Coal Mine Health and Safety Act of 1969 was enacted. Over time, the coal conveyer equipment being used and ventilation conditions underground have changed significantly, particularly the width of the conveyor belts and the use of belt air at the face. The effect of these new conditions on the effectiveness of suppression systems during a fire is not known. The National Institute of Occupational Safety and Health (NIOSH) conducted a study to evaluate the effect of air velocity, water sprinkler activation temperature and a limited water application time on the effectiveness of water sprinkler fire suppression systems to extinguish large-scale conveyor belt fires. The fire tests were conducted using both new and used 1.8-m-wide, fire-resistant rubber belt that met the Mine Safety and Health Administration (MSHA) flame-resistant requirement specified in Title 30, Code of Federal Regulations, part 18, section 18.65 (also known as the 2G test). Tests were conducted using standard response sprinklers with activation temperatures of 68° C and 141° C at air velocities of 0.5 and 5.1 m/s. Two tests were conducted at both air velocities for each sprinkler activation temperature, one using a new belt and one with a used belt. The results showed that the suppression system was able to suppress the fires in ten minutes to the point that a miner could extinguish it with a fire hose. However, in several of the tests, the fire reestablished itself a few minutes after the sprinkler water supply was cut off and quickly grew out of control. This report discusses the large-scale experimental configuration, the installation specifications of the fire suppression system and the results and conclusions regarding the effect of air velocity, sprinkler activation temperature and limited water application on the suppression system performance.

Jan 1, 2010