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By Thomas Camus

"STATE OF THE ART The current situation of ground prediction in TBM is of great concern for tunnelling contractors, and the subject has been quite widely addressed over recent years, for instance by the French Eupalinos 2000 programme, which tested a variety of technological approaches. A number of interesting projects were executed, such as the SOUND and OnSITE projects, part of the German Geotechnologien programme: these projects are focused on seismics, while other techniques such as radar are not considered. Little information and even less industrial applications have been released apart from lab simulations and numerical modelling. A number of field tests in different configurations have been carried out for compression waves, but the results were often not reliable or convincing, and therefore little has been reported in open literature. The advances seem somewhat promising as far as hard rock tunnels are concerned, but ignore the problem raised by concrete lined tunnels such as those bored by a TBM in soft ground. In such tunnels, sending and receiving seismic waves reliably through the lining (and the grouting behind) is quite illusory. A careful analysis of the seismic prediction research results show that the main problem is the source which is poorly controlled and poorly coupled (for instance in concrete lined tunnels, or in approaches that use the TBM cutter head vibrations as a source). Generating waves perpendicular to the tunnel axis is also a problem: aligning with the tunnel axis makes more sense. Radar (Ground Penetrating Radar - GPR) has been used in geotechnical surveys for decades, and also for tunnel inspection. Attempts have been made at using it on a TBM (for example in Japan by Taisei) but mainly using off-the-shelf GPRs, which are not adapted to work in the specific environment of a TBM (because of its huge mass of metal, to start with). Very few results have been published, probably due to the somewhat disappointing performance of such implementations. At present, no TBM has a GPR incorporated in its cutter head that is able to predict ground characteristics while drilling. Another characteristic and drawback of most current methods is that they require dedicated periods of time during which other tunnel construction activities must be stopped. Due to the duration of such operations this leads to performing ground prediction on a weekly basis only, and therefore to a requirement for an investigation range of 100 metres approximately (in order to cover the TBM advance over the coming week). This is not achievable with boreholes made from the TBM, even less if a reasonable accuracy is required, as is usually the case."

Jan 1, 2016

By A. D. Pernichele

Four types of slope failures in waste dumps are recognized; debris flow, foundation failure, edge slump and blowout. The first three types, although at times causing considerable damage to property and interfering with mine operations, are not usually dangerous to personnel because of their slow rate of movement. The fourth type, the blowout, is considered to be the most dangerous because of its explosive mode of failure and lack of forewarning. The December failure at Verona dump is presented as an example of a blowout. Data from piezometer holes indicate that the blowout was due to perched leach solutions near the face of the dump. The level of water in this zone is very closely related to the rate of leach water application on the dump. A detailed analysis of core samples indicated that the impounded water was due to a zonal accumulation of Jarosite which is associated with a buried dump surface. Drain holes proved successful in controlling perched leach solutions in the critical zone near the face of the dump. Monitoring of the water level in the perched zone is considered essential in preventing blowouts. The use of drain and piezometer holes has allowed an increase in the rate of water application on the dump while preventing additional failure.

Jan 1, 1971

By T. Glenn Pobanz

For more than fifty years Walter Kidde & Company, Inc. has engaged in the development and manufacturing of both fire suppression and detection equipment. Our initial work for the mining industry centered around dry chemical extinguish-ing equipment for mine haulage cars. To fulfill this need, Kidde engineered and developed a manually activated dry chemical system to provide protection for the wheel wells, cable reels, and motor compartments on various pieces of underground mining equipment. The next phase of mining operation fire protection for Kidde was the area of detection and automatic fire suppression for conveyor belt heads. Early in 1970, a field evaluation program and belt head demonstration unit was taken to 10 areas throughout West Virginia, Kentucky, Pennsylvania, and Ohio. This extensive field program was a joint effort with National Mine Service Company of Pittsburgh, Pennsylvania. The objectives of this program were to study mine hazard problems and to determine further requirements of mine operators. The resulting belt head detection system centered around our Continuous Strip Thermistor Core heat detector in conjunction with a Kidde stored pressure dry chemical extinguishing system. Continuous Strip detection has been in use for a number of years in aerospace applications. It has proven flexible and reliable for the rapid detection of overheat conditions and fires in engines of over half the aircraft flying today. The use of Continuous Strip gave us flexibility to protect surface areas over and around the actual conveyor belt head equipment. The nature of Continuous Strip allows us to bend, wrap, and attach the 20 foot long thermistor element over the desired protected points of the belt head. An integral part of the conveyor belt head protection system is the control panel utilizing our comprehensive knowledge of fire alarm control logic. This panel contains supervision circuits which sound trouble signals upon detecting a break in the Continuous Strip; breaks in the wiring to the remote electrical extinguishing activators or to the alarm bell; or loss of primary power. A separate control module provides the adjustable alarm temperature setting of the

Jan 1, 1971

By T. P. Meloy

Oily and particulate befouled waste water discharges are a common industrial problem. Removing these water bourn emulsions from process discharges become increasingly important for both recycled and discharge water streams. Low cost, efficient and easily maintained systems are important. This paper evaluates design factor of such a water cleaning system -a trough filter. This is a continually operating filter that removes oil droplets by wires that wet the oil droplets.

Jan 1, 1994

By X. Yin

Surface force measurements based on atomic force microscopy (AFM) are used to describe the wettability of surfaces, specifically kaolinite basal plane surfaces at a high resolution. The kaolinite particles (~500 nm) were ordered with either the silica tetrahedral face exposed or the alumina octahedral face exposed, and surface force measurements were made between a hydrophobic diamond-like-carbon (DLC) tip and the kaolinite faces at pH 4, which is around the isoelectric point of the DLC tip. Under these conditions, the effect of electrostatic interaction should be eliminated and only van der Waals interactions and hydrophobic-hydrophobic interactions are expected to contribute to the surface force. The wettabilities of the surfaces were determined from the magnitude of the surface force. Wettabilities of the kaolinite basal planes thus determined were compared with the corresponding wettabilities of talc and mica surfaces, which are well-known to be naturally hydrophobic, and hydrophilic surfaces, respectively. The results reveal that the silica tetrahedral face of kaolinte has a modest level of hydrophobicity, which may be due to the lack of hydrogen bonding sites. In contrast, the alumina octahedral face showed no hydrophobicity, since a layer of surface hydroxyl groups is present at the alumina face, resulting in a relatively strong interaction between this surface and water molecules. It is expected that, with a better understanding of the surface properties of kaolinite, further improvements in flotation separations and dewatering may be achieved.

Feb 1, 2012

By John D. Smith

This paper describes how a knowledge of inherent rock characteristics can aid the mining engineer in his design of a mining method for a massive domelike structure. The paper shows how the pertinent rock characteristics were measured and analyzed and how the data obtained was used for design purposes. This analysis was conducted in 1964 in a producing salt dome, since that time, mine management has tried certain of the recommendations for mine design changes. The results of such changes are presented.

Jan 1, 1968

By C. A. Rowland

Comminution efficiency can be expressed in three different ways. Theoretical Efficiency: actual energy consumed, in a comminution machine compared to the theoretical energy required to make the size reduction. Operating Efficiency: compares the operating work index, calculated from production comminution machinery operating data, to the Bond Work Index from bench scale comminution tests or calculated from pilot plant operating data. Economic Efficiency: income compared to planned income. Using the Bond Work Index and operating work index, this paper explains how to determine operating efficiency.

Jan 1, 1998

By A. E. Aldrich

The Pittsburg & Midway Coal Mining Co. (P&M) is expanding a combination of surface and underground coal mining operation in Northeastern New Mexico. The York Canyon Complex, located 35 miles west of Raton, New Mexico, is aggressively pre-paring for increased production while meeting present requirements. Surface mining activities in four separate mining areas are preparing the complex for additional markets. Drag-line, truck/shovel, highwall miner and contour stripping are surface mining methods currently planned. In addition, P&M will move a Marion 8200 dragline from Kansas to the York Canyon site, finish constructing a $20 million coal preparation plant, and complete conveyor upgrades to the existing underground mine. This paper will describe the diverse York Canyon Complex.

Jan 1, 1993

By G. Blackwell

In the 1980's the Canadian surface mining industry was confronted by declining competitiveness in the face of low cost foreign producers and alternative materials. Low metal prices undermined profits for many mining companies. Restructuring, rationalization, capacity reduction, technology transfer, as well as favourable currency rates and higher commodity prices, helped revitalize the industry by the close of the decade. Now in the 1990's, however, the need to improve competitiveness for profitability has re-emerged in greater intensity. In the current climate it is considered that there are three strategies available to industry to ensure profitability: continued rationalization and restructuring; optimization of operating practice; and technology development and transfer.

Jan 1, 1993

By B. K. Baird

The application of cutoffs during and after pit optimization is examined for a lower-grade multi-metallic copper deposit. Pit optimization requires calculation of the net revenue associated with each mining block, incorporating parameters such as metal prices, grades and expected recoveries, mining and processing costs, smelting and refining charges, and penalty element and royalty charges. Sensitivity analysis during pit optimization reveals which of these parameters have a significant effect on the total contained resource tonnage and total contained non-discounted revenue in the optimized pit. The cumulative effect of changing several of the more significant parameters is investigated. Application of a copper cutoff grade to a block inside the optimized pit could result in the less than optimal extraction of the resource. A block with copper grade below the applied cutoff could be economic due to the presence of other metals. Conversely, a higher copper grade block may be sub-economic. The relative merits of applying a copper grade cutoff, "equivalent grade" cutoff and dollar value per ton cutoff to blocks inside the optimum pit are discussed.

Jan 1, 1999

By Robert M. Stevens

INTRODUCTION Truck haulage of muck in underground mining operations has been in use around the world for many years. From the late 1.950s into the early 1970s, haulage by means of load-haul-dump (LHD) units challenged haulage by means of trucks, and, to a large degree, LHD units dominated the trackless concept of moving mate¬rial underground. However, beginning in 1975, the emphasis began to shift back to the use of trucks, with LHD units assuming the secondary role of loading the trucks. Over the years, several manufacturers have been offering trucks with capacities ranging from 4.5 to 45.4 t (5 to 50 st) and with features that specifically address the constraints of underground operation. TRUCK TYPES Three different types of trucks are available for use underground, namely the tip dumpers, telescoping dumpers, and push-plate dumpers. The following para¬graphs describe each of these three types. Tip Dumpers Tip dumpers are designed to lift the truck box, using gravity to dump the load out the rear of the box. The primary application of these trucks is in the haulage of muck to the surface. However, these trucks also may be utilized underground, if a chamber is provided to accommodate the raised truck box at the dumping point. Fig. 1 illustrates a four-wheel drive truck with a 10.0-t (11-st) capacity; this truck is only 1.9 m (6.25 ft) wide and is designed specifically to follow narrow and winding vein courses. Fig. 2 illustrates a two-wheel drive tip dumper with a 24.5-t (27-st) capacity; this truck operates on 12% grades with firm dry surfaces and in modes both loaded-up and loaded¬down. Telescoping Dumpers Telescoping dumpers specifically address the constraints imposed by low back heights, and they represent an optimization of compressed envelope size with maximum payload capacity. Figs. 3 and 4 illus¬trate a typical telescoping dumper with a capacity of 18.1 t (20 st). Fig. 3 shows the box in position for carrying the load, and Fig. 4 shows the same truck with the box in the full-dumping position. Fig. 5 illustrates the loading and dumping cycles. Fig. 6 illustrates an approximation of how an LHD unit with an eject-o-dump (EOD) bucket would be used to load over the rear of this type of truck. Fig. 6 shows a low-profile telescoping truck in the first stage of loading, with a 6.8-t (7.5-st) LHD unit equipped with an EOD bucket. A loader with a smaller capacity and a narrower bucket would allow the bucket to fit inside the truck box and would achieve an even lower loading height than that illustrated. Gen erally, an LHD unit equipped with an EOD bucket and teamed with a telescoping dump truck can be used with a lower back height than could any other front-end loader and truck team. Push-Plate Dumpers Push-plate dumpers are similar to telescoping Bump¬ers, except that they do not have the two-stage box. Instead, push-plate dumpers utilize a single push plate to discharge the load out the rear of the box. Gen erally, these units are limited to payload capacities of about 9 t (10 st). Fig. 7 illustrates a typical push-plate dumper.

Jan 1, 1982

By Dennis Z. Mraz, Maurice B. Dusseault

INTRODUCTION In 1982 and 1983, as part of the mine design program for a New Brunswick potash mine, series of laboratory tests were conducted on potash and salt samples obtained during the mine exploration program. Concurrently, two rock mechanics monitoring programs were initiated in-situ the short term test at site A, and the long term test at site B (figures 1 and 2). SHORT TERM TEST (FIGURE 1) The purpose of this test was to obtain a quick confirmation of laboratory data and comparison of the rock response to excavation with the data available from Saskatchewan potash mines. Instrumentation in the main drift was installed on Day 4 after excavation, and the pressure cells were pressurized on Day 33. The packer type cells in the east crosscut were pressurized on Day 25. Excavation of the west cross-cut was begun on Day 71. This "mine by" test has provided important information required for calibration of laboratory tests, as will be shown later. The test data recorded are shown in a series of diagrams: a. initial room closures, main drift-figure 3; b. convergence graphs, main drift-figure 4; c. stresses recorded, main drift-figures 5 to 8 and, d. stresses recorded, east crosscut - figure15 Three types of instruments were used in this test. In the main drift. Saskatchewan Research Council water type hydraulic inclusions. encapsulated in potash, were used to measure lateral and vertical stresses. In the east crosscut a home-made packer type hydraulic inclusion was used to record radial stresses. The convergence stations consisted of end-grouted bars of reinforcing steel. LONG TERM TEST (FIGURE 2) The purpose of the long term test was to study the steady-state behaviour of a single opening in potash. Site B was placed in a remote area of the mine, well away from any anticipated early mining activities. To assure homogeneity of rock mass, the site was placed within a 30 meter thick vertical limb of a syncline-shaped ore body contained within a 360 metre thick vertical section of salt rocks. It was excavated mechanically as close as possible to the designed square cross-section for the entire length of the room. The square cross- section was chosen in order to achieve a comparable geometry in vertical and horizontal directions; that is, to eliminate any undesired effects of geometry interaction with lithostratigraphy which might have generated non-homogeneous behaviour. The site was extensively instrumented. As in site A, the convergence stations consisted of end-grouted bars of reinforcing steel. However, in order to avoid any measurement inaccuracies which would be associated with the brittle yield zone in the opening walls, bars were placed to the same depth in opposite sides of the opening. This allowed the convergence measurement to be obtained as the absolute convergence between two points, rather than the convergence of a single point with respect to a potentially unstable reference. For stress measurements the Saskatchewan Research Council cells were placed vertically, laterally and radially. Unfortunately, due to cost considerations, it was decided to use a compacted sand, rather than grout, to encapsulate the cells in the boreholes. Due to the extremely small strain rates, only radial ly placed cells could be adequately pressurized. The rest of the cells did not perform adequately, and are currently being replaced.

Jan 1, 1986

By J. R. Frederick

Introduction In response to industry needs and requests, Joy Mining Machinery is aggressively evolving a "game changing" system of underground development mining machinery. This entry-development system (EDS) project is concentrated on providing clean, protected and ergonomic operator areas and a quantum leap in productivity. To meet these goals, a virtually continuous and automated mining system is needed. Operators must be able to efficiently and ergonomically perform required functions from a protected environment. Health-related issues, such as respirable dust and noise, must also be addressed. A new concept system of integrated face equipment, from the continuous miner to and including advancing the section belt and services, is needed. Phase I of the project was aimed at developing the bolting technology required. Through acquisition of Cram Australia and joint development efforts with Canyon Fuel Co., Joy now has this capability. The test platform, known as the Joy Multibolter, has successfully completed its trials at Canyon Fuel's SUFCo Mine (Fig. 1). These tests validated the bolting technology and operator protection and ergonomic issues. Phase II includes the development and testing of an integrated mining machine (IMM) capable of advancing a heading 1.5 m (5 ft) while simultaneously cutting a fullface width (single pass) and installing a complete bolt pattern. The IMM, unveiled at MINExpo 2000, incorporates the bolting technology of the Multibolter with automated full-face cutting. This machine was designed to meet Joy's stringent health, operator protection and 300m/d (1,000-ft/d) productivity goals. Phases III through V will include development of processes and integrated haulage, ventilation, power, material handling and section services advancing equipment. Processes and equipment developed will be designed for seamless integration into a complete entry development system. The focus of this paper is the new continuous miner (IMM) development, proposed testing and expectations for the EDS. The IMM is scheduled for commissioning and trials beginning in April 2001 at the SUFCo Mine. SUFCo Mine Canyon Fuel Co. is jointly owned by Arch Western Resources and the Itochu Corp. of Japan. It owns and operates three coal mines in Utah: the SUFCo Mine in Sevier County, the Skyline Mine in Carbon and Emery Counties and the Dougout Canyon Mine in Carbon County. Incorporating state-of-the-art-mining technology and advanced mine management practices, the SUFCo Mine is one of the most efficient underground coal mines in the world. The SUFCo Mine is located approximately 50-km (30 miles) northeast of Salina, UT, in the Wasatch Plateau portion of the Uinta coal basin. The mine property consists of 9,900 hm2 (24,470 acres) with approximately 118 Mt (130 million st) of recoverable reserves. Public utilities and industrial concerns purchase approximately 95% of the coal mined at SUFCo through short- or long-term contracts. Coal reserves are extracted by longwall mining, and the panels are developed using continuous miners. The longwall face at SUFCo Mine is 283-m (930-ft) wide, and the seam height ranges from 2.4 to 4 m (8 to 13 ft). The two development sections are in full operation using Joy 12CM12 continuous miners. Two super longwall panels, under development in the northeast portion of the lease block, are 4,270-m (14,000-ft) long with a 283-m (930-ft) wide longwall face. Each panel contains approximately 5.4 Mt (6 million st) of recoverable coal. SUFCo has been a world leader in productivity for a

Jan 1, 2003

By D. Kelley

Due to pricing volatility and lack of availability of metals, the use of non-metallic materials of construction in metal extraction and refining processes has increased significantly in recent years Even though costs and availability of metals has improved recently, these factors are not expected to continue into the future as metal supplies once again fall short of global demand. One of the most common non-metallic materials of construction used in metal extraction and refining processes today is Fiberglass- Reinforced Plastic (FRP). These processes often require materials of construction that can resist acids and high concentrations of acid chloride salts. As shown in Table 1, FRP based on epoxy vinyl ester resin has the same chemical resistance to acid and better resistance to acid chloride salts than alloy C-276 and is superior to 2205 stainless steel in both cases.

Jan 1, 2010

BP Minerals America (BPMA) has completed its $400 million modernization of Kennecott Utah Copper's Bingham Canyon mine. Modernized facilities include the in-pit crusher, conveyor, ore storage area, concentrator, ore pipe-line, and tailings pipeline. With modernization, Kennecott will achieve substantial cost reductions. Bingham Canyon is now one of the lowest cost copper producers in the world. And the mine will continue to provide jobs and income in Utah for another 30 years or more. The Modernization Project began in February 1986. The work took 33 months. In late 1987, the construction force peaked at 1100 workers. Kennecott Utah Copper now employs a permanent work force of 2100 employees. Less equipment and more efficiency The modernization work has increased the scale of the facilities that transport and process copper ore. A large, continuous transport system has replaced multiple haulage units. Fewer but larger pieces of equipment have replaced many crushing, grinding, and flotation units. The result is less maintenance, fewer employees, and improved metal recoveries. About one-third of the $400 million modernization project is directed at improving ore movement by eliminating rail transport. The new concentrator accounts for the remaining two-thirds of the modernization investment. The modernized plant was started up in stages. In February, ore was fed to the crusher. By the end of March, the three grinding lines were operating and the flotation circuit was producing copper concentrate. In July, the design through-put rate of 70 kt/d (77,000 stpd) was reached. October saw the achievement of targeted metal recoveries that corresponded to the fully modernized operation. It is estimated that Bingham Canyon will annually produce 181 kt (200,000 st) of refined copper, 9.3 t (300,000 oz) of gold, 71 t (2.3 million oz) of silver, and 5.4 kt (12 million lb) of molybdenite. Current Bingham Canyon production is about 97 kt/d (107,000 stpd) and a similar amount of overburden. Some 70 kt/d (77,000 stpd) of ore are processed through the in-pit crusher and 27 kt/d (30,000 stpd) of ore are transported by rail. The Bingham Canyon mine has yielded more than 11 Mt (12 million st) of refined copper. In 1863, precious metals - silver-bearing lead - were first discovered in Bingham Canyon. What was once a mountain is now a huge open pit. In fact, the mine is the world's largest man made excavation. The pit is some 4 km (2.5 miles) wide and 0.8 km (0.5 miles) deep. Since the pit began operating in 1906, more than 4.5 Gt (5 billion st) of material have been removed. Modernized facilities Electric 21 m3 (27 cu yd) or 23 m3 (30 cu yd) shovels are used to load the haulage trucks. Both 154 t (170 st) and 181 t (200 st) trucks are used to haul the ore to a mobile, in-pit crusher. The crusher replaces three crushers located at the old concentrators. There are two dumping locations at the new crusher, which can handle two trucks a minute. The 1.5 x 2.7 m (60 x 109 in.), 745 kW (1000 hp) Allis-Chalmers gyratory crusher reduces the ore to 254 mm (10 in.) size or less and delivers it to the conveyor. The crusher and conveyor systems are capable of handling 9 kt/h (10,000 stph) of ore.

Jan 11, 1988

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

It is imperative that the hydrogeologic frame- work and the ground water flow systems of a Potential mine area be investigated prior to initiation of mining. Such investigations are necessary so that potential mine water inflow problems can be anticipated and controlled. Obvious difficulties can arise if large inflows of ground water occur in underground workings. Costs of water removal and disposal, if not foreseen and included in economic planning, can severely limit or eliminate profit from a mining operation. Safety problems associated with large inflows are an additional potential problem. The inflow of water into a mine necessarily alters the associated ground water system. The inflow also may impact nearby surface water regimes, depending upon the interaction of the surface water and ground water systems. The inflow of water into a mine inevitably causes a decline in the ground water levels in the vicinity of the mine. This decline may be restricted to a small area or it may extend a considerable distance, depending upon the hydrogeologic properties of the associated hydrostratigraphic units. A decrease in surface flow will result if the area of ground water level decline intersects a stream or spring that is hydraulically connected to the ground water system. Mine inflow problems historically have been evaluated after-the-fact. Premining hydrologic investigations only recently have been initiated. Brown's (1977)* report on anticipated hydrologic impacts of oil shale mining in the Piceance Basin of Colorado is an example of premining hydrologic planning. Observed difficulties inherent in designing a data collection network and then utilizing the data to predict mine inflow have led us to the preparation of this manuscript. This manuscript reflects a two-phase effort. The first phase describes and develops analytical and numerical techniques for mine inflow prediction. Emphasis in this section is placed on the selection and application of data collection techniques and data analysis procedures. The second phase applies the prediction technique(s) developed during phase 1 to two existing mines. The objective of phase 1 efforts is to evaluate the hydrologic variables associated with surface water systems and ground water systems that can affect the inflow of water into an underground mine. Subsequently these variables are incorporated into predictive techniques for estimating underground mine water inflow and its effect on associated ground water and surface water systems. The specific objectives of phases 1 and 2 are to: 1) Determine the hydrologic parameters required for predicting underground mine water in- flow and the potential effects of the inflow on surface and ground water systems. The hydrologic parameters represent both premining and past initiation conditions. 2) Outline methods for obtaining the necessary input data for the predictive techniques in pre- mining and post initiation periods. 3) Present existing numerical techniques for predicting underground mine water inflow including assumptions Ad required input. 4) Evaluate existing numerical techniques previously tabulated based on their applicability to mine water inflow prediction. 5) Attempt to develop analytical techniques for predicting underground mine water inflow. 6) Develop, through modification of existing models, a general mathematical model as a numerical technique for predicting underground mine water inflow and the resultant effects on surface and ground water systems. 7) Apply the predictive techniques to two existing mines selected for this purpose. 8) Compare the predicted inflow rates to measured inflow rates at both mines. 9) Compare predicted impact on district hydrology to observed impacts where possible. 10) Evaluate the predictive technique.

Jan 1, 1986

By M. P. Adamson, Bryson D. Trexler, R. V. Ramani, Mark L. Koenig, V. Rajaram, Alan Burton, C. S. Crocker, Kim. Y. C., A. G. Law, William C. Larson

Another record year vas experienced by the mining industry as the value of nonfuel minerals production soared to $23.5 billion-up 15% from the previous year. Based on preliminary estimates by the Bureau of Mines, US mine output was the basis for about $225 billion worth of minerals-derived materials. Metal mining accounted for $8.5 billion, or 36%, of the total. Of the 22 metals produced in the US, 11 showed quantity increases, while 16 had value increases. Business Week's survey of 1200 companies (Mar. 17, 1980, pg. 81) revealed that 21 metals and mining companies had sales increases of 30% with earnings up 104%. Higher prices and increased demand for metals were primary factors behind the improved performance. The steel sector reported a 30% decrease in earnings for the year, but the slump was primarily the result of a $383-million loss for US Steel coupled with a 17% decrease in earnings at Inland Steel. Other firms did well in 1979, with earnings up 279% at Interlake and 269% at Kaiser Steel. Major obstacles facing the steel industry included substantial tonnages of dumped and subsidized imports, environmental expenditures, runaway inflation, and price controls. Turning to nonmetallic production, of the 49 commodities reported by the US Bureau of Mines, 29 had higher outputs and 39 had increased values. While production statistics may seem encouraging, government regulations continued to stifle industry efforts to develop needed new capacity. A case in point may be publication of the Environmental Protection Agency's simplified procedure to streamline five permit programs by consolidating them into one. The five included: hazardous waste management, national pollution discharge elimination, dredge and fill, underground injection control, and prevention of significant deterioration. The new procedure, published by EPA in the Federal Register, took 149 pages. AMC's J.A. Overton, Jr., noted that (MCJ, Nov. 1979, p. 39) to visually depict the complexity of the proposed regulations, Sell Oil Co. developed a flow chart mapping a path through one section. The chart required 10 days to map and two weeks for two draftsmen, using computerized drafting techniques, to plot the results. Overton said, "Perhaps the greatest service EPA has performed in preparing its proposal is to demonstrate again that federal regulations have become a complexity compounded by confu¬sion that culminates in a conundrum." While regulatory agencies promulgate new regulations, the lack of a cohesive government minerals policy remained a matter of significant importance to mine management. A non-fuel minerals study to provide input to the Administration for a national minerals policy program was a dismal shock, according to Rep. Jim D. Santini (D-NV). He said the first phase of the program, which was to identify industry problems, failed to come to grips with the hard decisions necessary regarding these problems, their impact, and their interrelationship. Santini said the report failed to acknowledge government's role in the process and absolved the government of any adverse influence on the domestic minerals industry. In response to the phase one report, eight senior mining executives testified before Santini's House Subcommittee on Mines and Mining concerning mineral supply, access to foreign sources, environmental protection, conservation and recycling, national security, and capital formation. By year's end, Rep. Santini promised to prepare a hard-hitting committee report on problems affecting the non-fuel minerals industry to provide a better base for subsequent policy recommendations (MCJ, Jan. 1980, p. 34). On the technological front, interest again centered on mechanization, automation, and scaleup to improve and expand production. Computer use increased for calculation of ore reserves, mine planning, operations research, and mineral processing.

Jan 5, 1980

By John W. Carson, David S. Dick

The last few years have produced new methods of characterizing flow properties of bulk solids. Abrasive wear and particle attrition can now be accurately measured in a laboratory and scaled up to field condi¬tions. In addition, the effect of large par¬ticles on a material's flow properties (cohe¬sive strength, bulk density, wall friction) can be determined as well as the effects of var¬ious additives such as freeze conditioning agents. New developments have occurred with devices to improve the reliability of solids flow. Much attention has been paid to find¬ing ways, by using inserts, to provide a mass flow pattern with relatively shallow hopper walls. Such devices can also provide reli¬able in-bin gravity blending of bulk solids or completely nonsegregated discharge. Other developments have occurred in the area of belt-to-belt transfer chutes and sealing screws. The technology and hardware for running bench scale flow tests on bulk solids samples and transforming the results into reliable designs has been in existence for nearly 30 years (Jenike, 1964; Marinelli and Carson, 1986). It has been shown repeatedly that reliable materials handling systems can be designed to avoid the problems of no-flow due to arching or rat-holing, erratic flow, segregation, loss of capacity, or spontaneous combustion. Running flow properties tests and carefully designing a bulk handling system are not expensive, particularly when compared to downtime costs and continual flow problems. New developments in characterizing flow properties of bulk solids include: • a shear tester for measuring the properties of run-of-mine ores containing large par¬ticles; • attrition testers to simulate a wide range of flow conditions; • a tester to measure the wear rate of liners in bins, feeder parts, etc.; • a dustiness tester to measure the potential for dust and the effect of dust control meas¬ures; and • techniques for measuring the strength of frozen ores and the effect of freeze condi¬tioning agents. In the area of improving the flow of bulk solids, there are new techniques for ensuring reliable flow without using up excessive headroom. These include: • belt-to-belt transfer chutes to handle po¬tentially sticky, abrasive, and dusty ores; • designs for the interface between a bin/ hopper and apron or belt feeders to avoid hangups and minimize feeder loads; methods to feed solids reliably into a pres¬surized process; and designs for handling systems that will mini¬ mize attrition, avoid segregation, or auto¬matically blend materials together. Characterizing flow properties Large shear cell It has long been known that the cohesive strength of a bulk solid is due mainly to its fine fraction. However, the effect of the larger particles (that were assumed to re¬duce the strength) could not be measured. The development of larger scale (205 mm¬diam and 0.6 x 0.9 m or 8 in.-diam and 2 x 3 ft) shear testers has made it possible to measure the properties of a wider range of materials. Now cohesive strength measurements can be made with up to 100 mm (4 in.) particles. The results obtained from the large shear testers, together with the results of the nor¬mal shear tests on the fine fraction, are used in the design of reliable stockpile reclaim systems and crusher feed systems. A high percentage of large particles can reduce the measured shear strength of a bulk solid. However, segregation may alter the size distribution and should therefore be considered (Carson, Royal, and Goodwill, 1986). In addition, larger particles will

Jan 1, 1989

By D. M. Range

Exxon is reclaiming an 80-hm2 (200¬acre) Wyoming uranium tailings basin in accordance with US Nuclear Regula¬tory Commission (NRC) requirements. The reclamation plan is designed to control the tailings and their radioactive emissions for at least 1000 years. Rigorous modeling and analyses of potential erosion by wind and water were used in planning. The use of riprap was minimized and the need for long¬term reclamation maintenance was eliminated. The resulting design is simi¬lar to regions stable from erosion cre¬ated by nature. The concepts used in the design and implementation have wider reclamation application. Highland history Exxon operated the Highland uranium operations from 1972 until mid-1984. Open pit and underground operations mined the roll front, sandstone ore body. The 2.5 kt/d (3000 stpd) mill employed conventional crushing, grinding, sulfu¬ric acid leach, CCD, SX, precipitation, and drying circuits. The mill produced 10 Mt (11 million st) of tailings. An¬other operator began commercial scale in situ leach operations in 1988. The tailings basin was constructed by deposition of mill tailings in a natural ravine dammed with a 30-m (100-ft) high by 1050-m (3400-ft) long clay core earthen dam built entirely downstream of tailings. Figure 1 shows the layout of the tail¬ings basin in 1984. Tailings were de¬posited along the dam face on the east side of the basin as well as in the west end and along the south edge until the early 1980s. This left a deep water pond in the middle of the basin. The deep pond was largely filled with tailings during the last four years of mill opera¬tions by the construction of two tailings causeways and the operation of a float¬ing tailings line. The pond water volume was mini¬mized by recycling tailings water to the mill and by operating a 32-L/s (500¬gpm) evaporation enhancement spray system during the last few years of mill operations. Regulatory requirements Several conceptual tailings basin rec¬lamation plans were prepared during the mine life. In September 1984, Exxon began preparation of a detailed plan in response to an NRC request. Water, Waste and Land Inc. (WWL) of Fort Collins, CO was selected to help pre¬pare the plan that was submitted to the NRC in December 1984. The design had to meet the require¬ments found in NRC regulations (NRC, 1985). These were based on regulations prepared by the Environmental Protec¬tion Agency (EPA, 1983) at the man¬date of federal law (UMTRA, 1978). The NRC regulations contain strict primary standards and a somewhat more flexible set of secondary standards. The primary standards include requiring that: • radiation from tailings be reduced to background levels; • radon gas emitted by the tailings be reduced to 20 pCi/m2/s from the surface of the reclaimed tailings pile; • tailings not be dispersed into the environment; and • these conditions endure without routine maintenance for at least 200 years and for 1000 years to the extent reasonably attainable. The lengthy time requirements pre¬sented a large engineering challenge. Radon is a radioactive, chemically inert gas with a 3.8 day half-life. It is released from uranium tailings by the radioactive decay of radium, which, in turn, is a radioactive decay product of uranium. Radon emissions from ura¬nium tailings can be controlled with an earthen cover thick enough that the radon gas decays to its solid, short-lived radio¬active daughters before the radon dif¬fuses through the cover material into the air.

Jan 1, 1989

By P. N. Eksteen, G. E. Greig

The South African mining industry currently installs about 6 km per month of plastic pipes for conveying compressed air and water. Plastic pipes have a number of desirable properties and are favoured by many mining engineers because of their low mass, which facilitates ease of erection and assembly, and their excellent corrosion resistance and durability. However, they are combustible, and if involved in an underground fire can contribute to the overall toxicity of the combustion gases emitted. Furthermore, plastic pipes have the potential to conduct a fire from one part of a mine to another, and, by so doing, exacerbate an already hazardous situation. The limitations of small scale tests and the need to establish a large facility where underground conditions could be simulated was recognized by the South African mining industry. This paper briefly describes the Industry's full scale fire test facility and presents the conclusions drawn as a result of a number of full scale combustibility tests conducted on pipes manufactured from unplasticized Poly Vinyl Chloride (uPVC) and High Density Polyethylene (HDPE). Based on these findings, some practical solutions to prevent the spread of fire along the length of pipe columns of various pipe diameters and suspension configurations are presented. Recently some new and exciting developments in the field of intumescent coatings have occurred which are proving effective in minimizing the incidence of fire underground on South African mines. The potential benefits of these coatings are briefly discussed. INTRODUCTION Plastic pipes used underground on mines are predominantly in the form of protective cladding on insulated chilled water pipes and as service pipes for the conveying of water and compressed air to the workings. Plastic pipes have found wide use in the aggressive environmental conditions encountered underground due to their favourable properties relative to steel; these include being lightweight, easily installed and virtually immune to corrosion. However, plastic pipes are inherently combustible and can contribute to the overall hazard in three ways: (i) they are easily ignitable and can provide the initial fuel for a fire. (ii) once ignited, they can propagate a fire along their length and by so doing ignite other combustibles in their path, thereby contributing significantly to the overall conflagration. (iii) they can contribute to the overall heat, smoke and toxic gas emissions from a fire. While conscious of the value of small scale fire tests, the South African mining industry recognized the need for full scale testing to determine the combustion characteristics of plastic pipes under conditions closely approximating those underground and, as a consequence, a full scale fire tunnel test facility was commissioned with the objective of being able to (a), accurately evaluate the potential hazard of the various types of the plastic pipes found in mines, (b) use this information to suggest suitable fire engineering practices to minimize the hazard and (c), to practically verify the efficacy of the practices suggested. THE FIRE TUNNEL The fire tunnel is situated on a mine in the Western Transvaal and is 40 m long with a cross section of 3 m x 3 m which is typical of many haulages currently in use in South African gold mines. The ventilation drawn through the tunnel is provided by a 110 kW centrifugal exhaust fan and air velocities can be varied from 0 to a maximum of 5,8 m/s.

Jan 1, 1993