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By Richard M. Michaels

In considering the use of underground space for civil works functions, there are several criteria that must be met if transfer of any of them is to occur. It is important to recognize that these criteria involve social and political as well as conventional engineering and economic considerations. It is the purpose of this report to identify certain of these criteria and develop a set of measures that will permit the evaluation of the potential benefit of transferring civil works functions underground. The work reported here was completed for the ASCE and is contained in the report, Undergound Space, 1973. If one examines the total direct and indirect economic return from the location of civil works in subsurface rather than surface space, sizeable benefits accrue. This occurs largely because design, construction, operating and maintenance costs of these systems are highly dependent upon the forces exerted on these systems by the surface enviroment itself. Transferring of these facilities to underground space would eliminate a large segment of those costs. It is equally true that civil works functions compete with other uses for surface space and generate significant conflicts with these competing uses. Highway construction takes land from other productive uses. Electrical transmission lines conflict in terms of safety and aesthetics with adjacent social activities. Assuming that construction costs can be made equivalent, the analyses in the previous report indicate that the locating of all civil works functions underground would yield an annual net savings to the society of the order of 60 billion dollars. It may be added that the benefits in terms of improved social perfromance would also be substantial. However, it is clear that the benefits from utilization of subsurface

Jan 1, 1974

By W. Z. Choi

In our previous work, a population balance model describing the simultaneous processes of size reduction and mineral liberation has been developed. The experimental verification of this model, however, has been limited to a binary mineral system containing only one composite particle class. When considering multiple classes of composite particles, the estimation of model parameters becomes much more difficult. In the present work, a procedure for estimating model parameters for a multiple composite class model has been developed. Emphasis is placed on the direct experimental determination of these parameters. The physical significance of the breakage rate function and the liberation function in characterizing the liberation process is also discussed.

Jan 1, 1987

By R. V. Ramani, C. J. Bise

INTRODUCTION Classification of coal seams as thin, medium, and thick is both subjective and relative and is more a reflection of the statistical mode of the thicknesses en- countered in a mining region. A geological factor more precisely understood as defined than thickness is the inclination of the seams to the horizontal, commonly referred to as the pitch or the dip of the seam. Thick and steeply pitching seams present far more challenges to the ingenuity of the mining men to devise methods for their extraction than the thick and less steeply inclined ones. In a marginal case, where a thick coal seam may be minable by surface or underground methods, the decision bears a close relationship to the technological developments in the mining scene. Near surface thick coal seams are a special gift of nature and are not as wide- spread as the ones that call for the application of under- ground methods. Recent advancements in surface mining equipment have made it possible to surface mine to great depths. Stripping ratios of 30: 1 (cu yd of over- burden removed per ton of coal) have been achieved. The extraction of thick seams which preclude the use of surface methods is the concern of this chapter. Specifically, the objective here is to outline the special technological factors in underground mining of thick coal seams and to illustrate through several case histories the methods that are practiced in the North American continent. For the purposes of this chapter, a thick seam is defined as one which is not readily minable with currently available equipment without modifications to the extraction method or the equipment. At the present time, a thickness of 4.8 m (16 ft) can be considered as the lower limit, although developments in equipment and methods can be expected to push this limit upwards. The data for the case histories were collected from two extended mine visits in 1975 and 1978, and from other published accounts of the mines visited. Reference is drawn particularly to reports by Bise, Ramani, and Stefanko (1976) and Rarnani and Kenzy (1979). THICK SEAM MINING METHODS The various methods of mining thick seams can generally be grouped into three major categories: full face methods, where the entire thickness of the seam is re- covered in a single pass; slicing methods, where the seam is extracted to its full thickness by horizontal or inclined slices either in an ascending or descending order; and caving methods, where the seam is undercut and the en- tire coal seam is recovered by caving the top coal into the excavated cut, or some variations of the basic sub- level caving method. As shown in Table 1, in each of these methods, the coal is removed by either room-and-pillar or longwall methods. Hydraulic mining and transport of the coal in open flumes or in pipelines are also practiced. Addition- ally, the void created by the removal of coal may be filled with a packing material (stowing) except in the caving methods. The pitch (dip) of the seam, particularly when it is greater than 0.26 rad (150), imposes severe operational limitations on mobile equipment. A more detailed general discussion of the methods can be found in Bise et al. (1976) and Vorobjev and Deshmukh (1964). Specific discussion on the many variations of the three basic methods can be found in the papers presented at the international symposia that were held in

Jan 1, 1982

By Jozef S. Stachulak

INTRODUCTION Conveyor belt fires represent a serious hazard in underground mines. This fact is well understood in most mining countries. The statistical data for Ontario mines (1) indicates that the majority of the conveyor fires occurred at the head pulley of the conveyor belt. The main emphasis for fire prevention and detection should then be concentrated at the drive pulley of the conveyor designed to transport non-combustible material. In the case of conveyor belt fires, it is of utmost importance to detect the hazard at an early stage, when the overheating has not yet reached the ignition point. Thermal noise detection sensors were installed in 1987 at several conveyor sites at Inco mines. PRODUCTS OF COMBUSTION THERMAL DEGRADATION OF CONVEYOR BELT Various investigations on conveyor belt fire hazards have been made in the past (2, 3, 4, 5, 6, 7). Several belt samples (6, 7) were subjected to thermal oxidative degradation. The belts used in these tests were neoprene, polyvinyl chloride and styrene-butadiene rubber (6). About 70 different compounds were identified in the volatile thermooxidative products. As expected, hydrogen chloride was the major toxic product formed, accompanied by relatively low quantities of other toxic compounds such as carbon dioxide, carbon monoxide, sulfur dioxide, carbon disulfide, hydrogen sulfide, benzene, chloroprene, cresols, chloroethanol, formic acid and C8 hydrocarbons. The research investigation (8) performed by the USBM indicated that "smoldering combustibles generate a very low level of CO compared to the level of smoke produced". The cited publication (8), states that, "During smoldering combustion, typical mine combustibles produce about 2.5 times more smoke than CO. As fires develop to their flaming stages, the trend reverses and about eight times as much CO is produced as smoke." The reader is referred to several significant sources of additional information pertaining to emission products that can be found in references (8), (9), and (10). ANALYSIS OF FIRE HAZARDS U.S.A. AND ENGLAND Nearly 15 percent of all underground coal mine fires in the United States involved ignition of conveyor belts (7). In many cases, the fires propagated through thousands of feet of belt entry and created losses in the millions of dollars. Some of these fires consisted of rubber belts. Some involved polyvinyl chloride and neoprene belts that were classified "fire resistant". Typically, a belt fire begins after the conveyor is stopped. The fire is initiated by burning coal dust and grease accumulation around defective rollers or belt drives that have become heated by friction (7). An analysis of 113 conveyor fires (11) showed that 41 originated at the driving pulleys, 32 at the defective rollers, and 40 elsewhere on the conveyor system. The severity of the fire, which can spread along a conveyor road, has been emphasized by the disaster at Creswell Colliery, Derbyshire, on September 26, 1950, when 80 men were overcome and killed by toxic fumes from the fire (12). The fire started from smoldering caused by friction, and developed into a flaming combustion, propagating along the roadway containing the rubber belt at an estimated speed of 40.2 m/s (90 mph), and travelling a total of 650 m (2100 ft) THE INFLAMMATION OF CONVEYOR BELTS THE MECHANISM OF IGNITION Defective operation of a conveyor belt may start a fire in several ways: - Friction of the belt on the drive pulley. - Jamming of a guide roller, and others.

Jan 1, 1991

By Duluth Engineers' Club, Engineers' Club of Northern Minnesota

"Prospecting and Exploration. Since the Lake Superior ores occur in pockets or distinct bodies and vary much as to character and location, the actual mining of the ores is preceded by much work of an exploratory character. This work includes prospecting and exploration.Prospecting is the term generally applied to the quest for surface indications of ore, or the conditions which would warrant the expectation of finding ore in the vicinity. It includes such 'operations as geological examination, dip needle work, shallow test-pitting, and trenching. The ore bodies of the Mesabi Range are non-magnetic, and dip needle prospecting is therefore valueless. On the Cuyuna range, however, magnetic attraction as evidenced by the dip needle has been extensively employed as a guide to the location of ore deposits; in other localities it has found limited application.Drill Exploration. After the presence of an ore deposit is known or suspected, resort is generally had to exploration by means of diamond or churn drills. On the old ranges geological conditions generally make this manner of ascertaining the exact limits of an ore deposit impracticable; so, if two or three adjacent drill holes develop considerable depths of ore, the sinking of a shaft for underground exploration, development and mining is generally considered warranted. On the Mesabi Range, however, the flat-lying and comparatively shallow characteristics of the ore formation warrant much more extensive drill explorations. On this range, then, an ore body 'is almost invariably followed· out with the drills, and its limits are determined to the point where the complete plan of development can be worked out in advance of actual mining operations"

Jan 1, 1920

By Stephen G. Hardcastle

3D-CANVENT is a software package that integrates advanced computer aided design (ACRD) true 3D graphics with a mine ventilation simulator. The package runs as a Windows(tm) application to access its printer drivers environment and does not need third party CAD software. It is composed of two primary modules: DMVENT and MINEDESIGNER. DMVENT is a traditional Fortran coded Hardy-Cross iterative ventilation network solver written in 1980 with thermodynamic capabilities. This module is relatively unchanged with the traditional data input options for branch type, specified or calculated resistances, fixed flows, and fixed or variable pressure fans. MINEDESIGNER is the graphics engine that optimizes the ventilation design process. It performs the front-end transformation of input data entered in the graphical interface into the correct format for the solver. At the back-end it reconverts the historically standard tabular data output from the solver into an easily viewed graphical format. ACRD features of MINEDESIGNER are used to generate a 3D wire-frame node and branch network of the mine's ventilation system. The network can be displayed in up to 4 views orientated to XYZ planes or a 3D view. All the views have zoom, pan, slice and rotate options. The graphical interface efficiently permits data entry and editing via a mouse with pick-and-point item selection. Branches can be found or added with "search" and "join" options. Visual interpretation is enhanced by the 16 colour options for branches and numerous graphical attributes. Network locations are readily identified by alpha-numeric names for branches, junctions and fans, and also the logical numbering of junctions. The program is also readily expandable for pollutant simulation and control/monitoring applications.

Jan 1, 1995

By Don S. McCoy, W. Joseoph Schlitt, Donald B. Swanson

Introduction Contract miners have long played an important role in the minerals industry. Historically, these were individuals who worked as underground miners and were paid by the ton or the day. These hardy types still ply their trade in the industry. However, today the term "contract miner" generally refers to an independent contractor who supplies the people and equipment needed to conduct surface mining operations for an owner or lease holder. At one time or another, contract miners have probably handled virtually every mineral commodity from sand, gravel, and stone quarrying to mining uranium and precious metal ores. In today's market, though, contract mining appears to be concentrated in just a few areas. These include precious metal mines, various types of solid fuel operations, and reclamation of abandoned mines. This article covers market areas, contract mining organizations and services, and their advantages and disadvantages. Gold and silver mines Although exact figures are not available, contract mining represents a significant element in the precious metals segment of US mining. For gold and silver, the growth in contract mining appears to be paralleling the overall growth in mining activity itself. Nevada is the largest gold producer in the US. The latest available data (1985) on open pit gold/ silver mines is summarized in Table 1 (Nevada Division of Mine Inspection, 1986). These figures show that contractors mined nearly half of the ore in 1985. Total tonnage mined, of course, was much higher for both owners and contractors. Assuming an average stripping ratio of 2.5 to 1, the total daily mining rate for the contractors was more than 145 kt (160,000 st). This translates into a contract mining rate of about 36 Mt/a (40 million stpy) based on 250 operating days a year. This approximate 50-50 split between owner and contractor also appears to hold for the future. Metals Economics Group (1987) published a survey of 51 short-range and 54 long-range precious metal projects being considered throughout the US. Contacts with owners, requests for proposals, and past corporate philosophies all suggest that about 55% of the open pit operations scheduled for the near-term will be contract mined. This represents an estimated 26 kt/d (28.5 stpd) of ore. A similar assessment of the long-term projects indicates that about 12 kt/d (13,000 stpd), or just under 50% of the open pit tonnage will be contract mined. There are several reasons why gold and silver represent such an attractive opportunity for contract miners. One is that many properties are owned or leased by individuals or partnerships that are essentially entrepreneurial. Many of these owners or lease holders have minimal mining experience. They must therefore either rely on contract mining organizations, sell an interest in the project to another mining company that will become the operator, or hire an experienced management team. For about the same reason, junior mining companies also represent good prospects for contracting. This is because most junior companies focus on exploration. They would prefer to spend their generally limited funds on drilling programs. Thus, they tend to minimize additional capital requirements by using contract miners or farming out the property to another mining company. Another good reason for using contract miners is the size of the deposit. Many properties are so small that their reserves will not support the capital expenditures

Jan 12, 1987

By D. A. Norrgran, J. A. Wernham

Introduction In view of the continuing depletion of natural resources, requirements for energy conservation, plant optimization and hazardous waste management, and environmental consciousness, recycling and secondary recovery of materials are being developed with increasing interest Many secondary recovery processes are being developed that apply mineral processing technology. Gravity separation, magnetic separation and flotation have all emerged as explicit methods for recovering residual values from various process streams and hazardous constituents from waste streams. Recently, a technique for concentrating metallics has been successfully reintroduced. The concept of the alternating magnetic field eddy-current separator was developed over a century ago using rudimentary electromagnets (Patent 400317 issued to Thomas Edison on March 26, 1889). Widespread acceptance was, however, hindered by high costs, marginal results and poor economic incentives for secondary recovery. With the recent increased interest in secondary recovery and the development of high-energy rare earth permanent magnets, the eddy-current separator has now gained prominent acceptance. The separation forces imparted by an eddy current separator are the function of the magnetic field intensity and the alternating frequency of the magnetic field. The evolution of permanent magnets has provided a cost effective alternative for the generation of high-intensity magnetic fields. High-energy rare earth permanent magnets have substantially reduced the capital and operating costs over the electromagnetic circuits employed in antecedent eddy-current separators. Specifically, in recent years, the strength of permanent magnets has increased several fold with neodymium-boron-iron rare earth magnets now providing an energy product of 35 million gauss-oersted. Figure 1 shows the evolution in the strength of permanent magnets. The development of these rare earth magnets has led to the design of circuits possessing a magnetic force an order of magnitude greater than that of conventional permanent magnetic circuits. The eddy-current separator has been successfully applied in several metal sorting and recovery operations. Most common is the sorting of metal from shredded automobile scrap and municipal waste. With the advent of high-frequency, high-intensity magnetic fields, the separator has advanced to the point where it has direct application in the beneficiation of fine sized metals. Relatively fine sized (-1/4 in.) metal bearing slags and spent foundry casting sands as well as precious metal hearing electronic scrap demonstrate excellent metallurgical response to the eddy-current separator, and applications have progressed to finer materials and more selective separations. Theory Electrical currents are induced in all conductors when exposed to an alternating magnetic field. The induced current generates a magnetic field in the conductor that opposes that of the alternating magnetic field. Figure 2, is a schematic illustration of generated eddy currents. In this particular example, the alternating magnetic field is produced by a series of permanent magnets mounted on the circumference of a rotor. The permanent magnets have alternating polarity. As the rotor revolves, an alternating magnetic field is produced, and the rate of revolution determines the frequency of the alternating magnetic field. When a conductor, such as a metallic disc in the illustration, is placed in the alternating magnetic field, a closed loop current flow occurs. The current loop in the conductor produces a magnetic field that is a mirror image of the alternating magnetic field. At any point in time, the induced magnetic field in the conductor directly opposes the alternating magnetic field. This opposition of magnetic fields produces an instantaneous repulsion in the conductor. The conductor is consequently repelled from the rotor. The alternating magnetic field has no effect on nonmetallic materials as they pass through the magnetic field and discharge the rotor in a natural trajectory. [ ]

Jan 1, 1992

By B. Beatty

Discussion by B. Beatty This is a response to M.J. Hrebar's paper, which tracks in detail the highly productive stripping capabilities of modern large bulldozers at two separate installations. It proves the old adage "economies of scale" and quantifies the advantage of being able to move large loads of dirt over comparatively short haul distances by sliding rather than carrying the load. This discussion examines whether a bulldozer is the most efficient way of using this geotechnical principle. Case 1 - Study eastern US: Contour coal stripping application The energy consumption for the Case 1 mining method (eastern US contour stripping operation) shows the D11 averaged 117 L (31 gal) of diesel fuel per scheduled hour, and produced 333.3 bank m3 (436 bank cu yd) per scheduled hour (0.351 L/m3). Assuming 1 L (0.26 gal) of diesel fuel contains 10.4 kW-hr of energy, the unit expenditure of fuel energy for Case 1 Study was 3.66 kWh/m3. Another way of shifting this material is by using an overburden slusher (OS), as shown in Fig. 1. The OS as shown in Fig. I. The OS is electrically powered and consists of a manned low wall winch (LW ), a remote controlled (radio telemetry) highwall winch (HW) and a hoe. The winches operate to pull the hoe back and forth across the pit. The hoe carries load in the forward direction ahead of a rear mounted blade and returns empty. A spreadsheet (Lotus 123) simulation package, BOSSIM, models the OS operation on a load by load basis over a wide range of pit configurations. The geometry shown in Fig. 2 of the paper was modeled using BOSSI M and the results are summarized in Table1. Table 1 shows a LW with 150 kW (200 hp) and a HW of 120 kW (160 hp) installed power, towing a 15-m3 (19-cu yd) hoe, would accomplish the Case 1 stripping task at about the same scheduled stripping rate (D11=333 m3/hr or 433 cu yd/ hour, OS=342 m3/hr or 447 cu yd/ hour). It was assumed the winches in this equipment have an electromechanical efficiency of 90%. This gives an average electrical energy expenditure rate of 151 kW-hr, and a unit energy expenditure level of 0.44 kWhr/m-3(0.57 cu yd/kWh)-an improvement of 8291%- over the D11 alternative. The energy comparison illustrates the importance of using more efficient electric power if possible, ahead of the internal combustion engine, and delivering power from a fixed ground position rather than through a moving track system. Another important energy observation is the comparative action of the dozer blade and the OS hoe. The former implement continuously digs and spills as it works the load forward. The OS hoe digs the load until full, the blade then lifts due to the contained load pressure and a shear plane forms under the mass, enabling it to slide forward in a force reduced - transporting mode. Economic comparison Table 2 operating cost per scheduled hour and unit cost comparison follows a format used in Tables 2 and 3 of Hrebar's paper and compares Case 1 Study with the OS alternative. Drill and blast costs are kept constant in both cases, although when larger OS units are considered, it is expected these costs will reduce because bigger hoe sizes will permit larger rocks to be transported efficiently.

Jan 1, 1995

By K. A. Smith, X. Cheng

Introduction The separation of minerals from complex sulfide ores by flotation requires careful control of the surface chemical behaviors of the minerals constituting the ores. When the minerals in the ore are finely disseminated, extremely fine grinding is needed to liberate the minerals. Fine grinding is not only an energy-intensive operation, but also influences the flotation behaviors of liberated minerals. Steel balls are commonly used for grinding. Galvanic interactions occur not only between sulfide minerals and grinding media but also among the different sulfide minerals themselves. The electrochemical characteristics of chalcopyrite and pyrrhotite when they are independently coupled with steel grinding media and the relevance of these characteristics to the flotation behavior of these minerals have been exten¬sively studied (Cheng and Iwasaki, 1992: Li and Iwasaki, 1992). These studies dealt with reactions in two electrode combinations, simulating mineral-grinding medium and mineral-mineral interactions. In practice, however, it is more common to involve more than one sulfide mineral. In a study of a pyrite-pyrrhotite-mild steel system, Pozzo et al. (1989) provided evidence that a mineral having an intermediate potential could behave differently, depending on the condi¬tion of galvanic contact. This investigation was undertaken to study a chalcopyrite-pyrrhotite-mild steel system to gain better in¬sight into the mechanism of multiple galvanic interactions. A three-electrode system was chosen to simulate a typical copper sulfide ore. The sulfide minerals were represented by chalcopyrite and pyrrhotite. The grinding medium was rep¬resented by mild steel. The nature of galvanic interactions involving these minerals and grinding medium in an oxygen atmosphere was investigated in a three-electrode combina¬tion, and an attempt was made to relate the laboratory flotation results of pyrrhotite-chalcopyrite ores. Possible implications to the differential flotation of complex sulfide ores are discussed. Key words: Electrochemistry, Flotation, Sulfide ores Experimental The pyrrhotite (57.74% Fe, 0.09% Cu, 4.84% Ni and 36.54% S) sample was obtained from Falconbridge Ltd.'s Strathcona mine, Ontario, Canada, and the chalcopyrite (26.00% Cu, 28.57% Fe, 0.05% Ni and 31.52% S) sample came from Transvaal, South Africa. Mild steel (AISI 1020, 0.2% C) was chosen to represent grinding media. A conventional three-electrode cell was used in the elec¬trochemical measurements. The working, counter and refer¬ence electrodes were, respectively, sulfide minerals or mild steel, graphite rod and saturated calomel electrode (SCE, 0.245 V vs. standard hydrogen electrode). The exposed areas of the sulfide minerals and grinding medium electrodes were approximately the same (0.17 to 0.19 cm2). The preparation methods of working electrodes, the reagents used, the experi¬mental procedures and the instrumentation were the same as those described previously (Cheng and Iwasaki, 1992). Rest potentials of pyrrhotite, chalcopyrite and mild steel were measured individually in 0.05-M sodium sulfate solu¬tions. Combination potentials and galvanic currents of the sulfide minerals and grinding medium were measured in short-circuited two- and three-electrode systems. All poten¬tials in this study are quoted in volts with respect to the standard hydrogen electrode (SHE). To simulate the constantly abrading conditions in a ball mill, rest potentials, combination potentials and galvanic currents were also measured by rotating a porcelain ball against chalcopyrite, pyrrhotite and mild steel electrodes in

Jan 1, 2000

Keeping an eye on the future while honoring the past, Frank McAllister, chief executive officer of Stillwater Mining Co., and life member of the National Mining Hall of Fame and Museum, gave the keynote address to a packed house during the 23rd National Mining Hall of Fame and Museum Banquet in Leadville, CO on Sept. 11. McAllister?s company made news just days before the induction ceremony, when it was announced that Stillwater Mining would acquire Marathon PGM Corp. for $118 million (ME, Oct. page 14).

Jan 1, 2010

By Albert H. Fay

The gypsum industry makes significant contribution to the wealth of our nation. This non-metallic mineral ranks in the highest order with America's most precious resources. Evidence of this is borne out by the fact that gypsum is a vitally necessary element in the building, industrial and agricultural industries. In 1967, the United States consumed 14 million tons of this valuable ore. Of .this amount 69% was required by the building industry for the manufacture of building materials. An additional 21% was directed to the industrial market, for the manufacture of Portland Cement and for other industrial uses. The remaining 10% was used for agricultural purposes.

Jan 1, 1969

By G. G. O. O. Uwadiale, M. A. Kollere, M. A. U. Nwoke

Birnin Gwari and Alawa graphite deposits were upgraded by flotation. Using 0.5 kg/t kerosene and 0.25 kg/t sodium silicate, flotation concentrates grading 90.7% and 88% Cg with recoveries of 86.2% and 84% were produced from Birnin Gwari and Alawa graphite samples, respectively. The combination of 0.5 kg/t kerosene, 0.2 kg/t pine oil and 0.25 kg/t sodium silicate yielded high-grade graphite (92.3% Cg) with a recovery of 83.9%. Flotation with refined rubber nut oil and kerosene, at a 1:1 ratio, produced a concentrate grade of 86.3% Cg with a recovery of 90.4%. With acid leaching of the flotation concentrate, a high-grade graphite concentrate containing 98% Cg was produced.

Jan 1, 1998

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

The strength of anchorage of various types of roof bolts in permafrost depends on the types of rock in which they are installed. Variation in the mine thermal regime results in weakening of the roof and walls, and at times total loss of ice cement bonds takes place in the surrounding rockmass. In mines with positive ventilation air temperature, the wedge or expansion types of bolts may become weakened due to the transfer of heat into the anchorage through the metal rod and subsequent thawing of rocks. Effective operation and evaluation of roof bolt systems for mine workings in permafrost require predictions of the dynamics of wall thawing by warm ventilation air. This paper presents an analysis of the temperature distribution in the frozen host ground surrounding roof bolts, the area of interest being the vicinity of the anchors for point anchored bolt and the entire length for split set bolt

Jan 1, 1994

By Andrew Chegwidden, Tara Wilk

Construction of Shaft 17B-1, a part of The New York City water supply system, required an excavation of approximately 38 m (123 ft) through water-bearing overburden soil and more than 152 m (500 ft) of gneiss bedrock to make a connection to Water Tunnel No. 3. Ground freezing was the specified method to provide temporary earth support and ground water control for the overburden material. The ground freezing process relied on a supplemental geotechnical investigation and comprehensive laboratory testing of the frozen soil to evaluate the potential long-term creep behavior of the soil and the elimination of a temporary lining system. This paper describes that process as well as the drilling and installation methods of the ground freezing system. Additionally, the methodical system of excavation and insulation is discussed that permitted a safe and timely excavation followed by the installation of the permanent concrete liner.

Jun 13, 2023

By Michael Michaylov

INTRODUCTUCTICN The coal seams at Bobov dol basin are highly prone to spontaneous combustion in Bulgaria. The risks of endogenous fires (SPONCOM) exist mainly in gob area. During the last 15 years three different technologies for preventing and fighting SPONCOM have been applied in underground mines of this basin: pulp from hard material and water, foam and inerting with nitrogen. On the basis of experimental application of these technologies as well as theoretical and experimental research a new technology - Foam Pulp Technology (FPT) was worked out. It is very effective in difficult to reach places in a gob. BACKGROUND AND THE NEED OF FPT Application of the three methods mentioned above showed advantages as well as some disadvantages, allowing determination of their effective application scope depending on geological and working conditions. Their advantages can be summarized as follows: - Foaming: high viscosity at low density, which allows forming a thick layer; good surfactant wetting properties; excellent effective cooling during evaporation in high temperature zones. - Pulping with mixture of solid particles (e.g., fly ash) and water from surface installation: easy to apply in a mine; low operational and energy costs due to gravity transport from the surface to the gob: good ecological results because the waste material is returned back underground. - Inerting with nitrogen: volumetric action which doesn't restrict the thickness of treated coal pile; easy to apply in a mine; controllable nitrogen delivery depending on fire hazard conditions. Main disadvantages of the same three methods are: - Foaming: rapid disintegration due to intrusion into heterogeneous and porous media and high costs due to the need of frequent large volume foam flows (short foam's life); - Pulping: inapplicable for coal piles with high thickness which is of great importance in the cases of seams with ununiform hypsometry (rapid and frequent changes in a seam configuration) and/or great volumes of coal or ore piles particularly heated to a high temperature; strong dependence on gob floor declination; water draining of the working face area which impedes the movement of supports and chain conveyor's. - Inerting: relatively high costs hardly depending on the rate of face advance (fig.1) in comparison with the other methods; it is difficult to ensure sufficient oxygen at the exit end of the face. This problem is crucial in faces with a high level of methane emissions; accelerating of the oxidation and spontaneous heating of the pile after turning off the nitrogen supply. Inerting should be applied again at higher capacity of nitrogen supply (fig.2) and what is more important, it shouldn't be

Jan 1, 1995

By M. Albrecht

I have been following the discussions on coal degradation started by the article "Degradation process in coal slurry pipe-lines," by M.G. Ayat and B.C. Scott. I agree with Mr. Dasher's comments in January that the title of the article was perhaps misleading as the information appeared to pertain more to processing the coal than to transportation. Specifically, the authors' conclude that the cyclone had an important impact on the degradation. Dr. Ferri's comments on degradation in other physical processes was very correct, handling of coal in a turbulent manner has an impact on the size distribution of the coal. This is one reason that when I design a coal system (mine to rail car) I try to minimize the total number of transfer and processing steps consistent with achieving the desired product specifications. Another area, mentioned in passing in the article, the impact of natural causes and the mining step, is of particular interest to me. A mine in West Virginia had a variation of 50% in the -6.35 mm (-0.25 in.) raw coal depending upon the direction of mining. Since it was a single unit mine, this significantly impacted the final product. Two articles on mining's impact on size distribution are: • Stefanko, R., Chopra, 1K., and Ramani, R.V., 1974, "Mining influence on size consistency and washability characteristics of coal," SME Fall Meeting, September; and • Albrecht, M., and Olsen, D., 1980, "Mining methods impact on coal preparation plant design," 1st International Mine Planning and Development Symposium, September. Both articles presented information on variation of size distribution and cleaning plant performance on a coal depending on the mining method used. Both articles considered a single coal seam at one mining operation. Stefanko's article considered continuous and conventional mining of coal in Pennsylvania. Albrecht's article considered surface, underground continuous, and underground hydraulic mining of a coal in British Columbia. Both concluded that mining methods do impact coal characteristics. I agree with Mr. Ayat's comments in reply to Dr. Ferri's discussion, "... degradation is a multi-faceted phenomenon involving many natural and man-made factors." As efforts are being made to produce a higher quality (lower sulfur) coal, this topic becomes even more important

Jan 1, 1989

By M. G. Ayat

I have been following the discussions on coal degradation started by the article "Degradation process in coal slurry pipe-lines," by M.G. Ayat and B.C. Scott. I agree with Mr. Dasher's comments in January that the title of the article was perhaps misleading as the information appeared to pertain more to processing the coal than to transportation. Specifically, the authors' conclude that the cyclone had an important impact on the degradation. Dr. Ferri's comments on degradation in other physical processes was very correct, handling of coal in a turbulent manner has an impact on the size distribution of the coal. This is one reason that when I design a coal system (mine to rail car) I try to minimize the total number of transfer and processing steps consistent with achieving the desired product specifications. Another area, mentioned in passing in the article, the impact of natural causes and the mining step, is of particular interest to me. A mine in West Virginia had a variation of 50% in the -6.35 mm (-0.25 in.) raw coal depending upon the direction of mining. Since it was a single unit mine, this significantly impacted the final product. Two articles on mining's impact on size distribution are: • Stefanko, R., Chopra, I. K., and Ramani, R. V., 1974, "Mining influence on size consistency and washability char¬acteristics of coal," SME Fall Meeting. September; and • Albrecht, M., and Olsen, D., 1980, "Mining methods impact on coal preparation plant design." 1st International Mine Planning and Development Symposium. September. Both articles presented information on variation of size distribution and cleaning plant performance on a coal depending on the mining method used. Both articles considered a single coal seam at one mining operation. Stefanko's article considered continuous and conventional mining of coal in Pennsylvania. Albrecht's article considered surface, underground continuous, and underground hydraulic mining of a coal in British Columbia. Both concluded that mining methods do impact coal characteristics. I agree with Mr. Ayat's comments in reply to Dr. Ferri's discussion, "... degradation is a multi-faceted phenomenon involving many natural and man-made factors." As efforts are being made to produce a higher quality (lower sulfur) coal, this topic becomes even more important

Jan 1, 1990

By I. R. Miller

The purpose of this discussion is to try to offer some suggestions as to how the zinc and lead markets will perform over the balance of the decade, review some of the factors influencing production and come to some conclusions regarding prices. The past 2 years have been simply awful in terms of consumption, prices and overall profitability for the industry. In looking ahead to the year 2,000, it is probably not a bad idea to look back about 20 years to see where we have been and how we got to where we are. Zinc consumption from 1950 to the early 1970's increased steadily and rather nicely at an annual average rate of close to 4.0%. It was starting in the early 1970's that things started to get volatile and at times, difficult for the industry. The general economic boom of the early 70's was marked by annual consumption rates close to 5.0 million tons, a level not seen again for over 10 years. The First Oil shock of 1973-74 was followed by the sharpest consumption drop in memory. In two years, Western World zinc consumption fell 27%. While our chart shows a rather dramatic recovery thereafter, the average annual growth rate since 1970 has only been a rather modest 1.5%.

Jan 1, 1992