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By Brian E. Davies, Bobby G. Wixson

INTRODUCTION The mining, extraction and refining of metals into useful materials is an honorable profession with an ancient and productive history. Civilizations have long depended upon metals for their implements, structures and machinery necessary for survival, growth, and development. Today, as in the future, metals and minerals are essential to the progress of mankind as we approach the 21st century with technologies and concepts that will require new concepts and production from the metal mining and processing industries. Today is the tomorrow we talked about yesterday and the implementation of new ideas is essential if we are to achieve "Frontier Technology in Mineral Processing". Unfortunately, the state-of-the-art mining, processing, smelting and refining methods historically employed were less than desirable environmentally, which has caused todays public to view the metal mining and processing industry with concern. Presently the ancient rock and chat tips connected with old Roman lead-zinc mines which were worked in Wales are causing stream contamination problems that must be corrected at considerable expense to the local governmental agencies (Firth, J.N.M., et al., 1981). However, improved progress has been made and the environmental control and clean-up procedures utilized by modern lead-zinc operations in the "Viburnum Trend" or "New Lead Belt" of South East Missouri may serve as a model to illustrate how alternative technologies, environmental concerns and cooperative efforts contribute to efficient mineral production within the confines of a U.S. National Forest region without significant environmental impacts. This area received international attention when it was opened within the boundaries of the Clark National Forest in the mid 1960's. By 1970 it ranked first in the world by producing 432,576 short tons of lead ore and has since continued in this position producing 451,153 short tons of lead ore in 1983 representing 91% of the total United States lead production (U.S. Bureau of idines, 1983). This production comes from a mining area approximately 40 miles long by 10 miles wide extending in a north-south direction. Geologically the lead is disseminated throughout the Cambrain age Bonneterre Formation, mostly a dolomite, at depths ranging from 700 to 1200 ft. (213.4 to 365.8 m). Galena (lead) is the principal ore mined with lesser quantities of sphalerite (zinc), chaleopyrite (copper) and silver being recovered as economic co-products. This new mining development in one of Missouri is most scenic areas could have led to a classic confrontation between industries and environmentalists. Instead, through the cooperation of industries, governmental agencies and university researchers, the Ozark environment has been protected, and may serve as a model for future coexistence of the mineral industry with the environment. Cooperative Program Development As early as 1965 five new mines and two lead smelters were being prepared for production and environmental concern was an essential part of that planning. All the mining companies became acquainted with representatives of the various federal and state agencies and scientists from the University of Missouri-Rolla (UMR). During the next six years, St. Joe Minerals Corporation, American Metal Climax, Inc., Cominco American Incorporated, Ozark Lead Company (Kennecott) and American Smelting and Refining Company participated in cooperative research efforts. The companies also participated in information exchange meetings, the development of improved analytical methods and the utilization of research findings in the modification or construction of pollution-abatement facilities. Early on research was carried out to determine background values and to establish natural baselines of trace metals in the surrounding environment. The unusual topography and drainage pattern of the mining area were helpful since it was possible to establish control sites and test sites below individual mine-mill operations to study waste-water effects on the receiving streams. The location of the ‘New Lead Belt', mines, mills,

Jan 1, 1985

By B. K. Sternberg

Since 1986, we have developed a series of electromagnetic (EM) sounding systems at the Laboratory for Advanced Subsurface Imaging (LASI). These may be categorized by their frequency ranges: low frequency (LF, 30 Hz to 30 kHz), medium frequency (MF, 1 kHz to 1 MHz), and high frequency (HF, 30 kHz to 30 MHz). We have applied these EM systems to a variety of targets, including: mapping of plumes of contaminated groundwater, tracing leachate for an in-situ leaching operation, locating buried hazardous waste, identifying subsidence-prone zones and detecting buried voids. These systems are also applicable to exploration for ore bodies. Currently, our sounding system uses a vertical magnetic field vector. Previous systems of this type required a slow and relatively inaccurate manual rotation of the sensing coils. The systems also used occasional calibrations of the measuring system. We have developed a 3-dimensional mathematical rotation procedure for the data, as well as a simultaneous, high-accuracy, calibration method. These new developments have greatly increased the speed of the measurements, as well as improving the accuracy of the ellipticity data. We use stepped-frequency measurements along with narrow-band filtering to allow accurate measurements, even in high-noise areas. We typically obtain ellipticity measurements at eleven frequencies and along densely sampled profile lines. We use neural networks to help interpret the measured ellipticity patterns. Neural networks can be trained to relate the measured ellipticity patterns to earth resistivities or to discrete targets buried in the earth. The neural network training may use theoretical EM responses calculated from computer codes for relatively simple structures. The theoretical EM response for many complex targets that are of interest cannot currently be reliably calculated, even with the largest supercomputers available. We have therefore chosen to do some of the training using theoretical EM responses generated in a full-scale modeling tank located at our Avra Valley Geophysical Test Site. We have also developed data-visualization software to help interpret the large data sets produced by this system.

Jan 1, 1997

By V. Králová

The TESCAN Integrated Mineral Analyzer (TIMA), an automated mineralogy system, enables fast and quantitative mineral analyses of rocks, ores, concentrates, tailings, leach residues or smelter products. It combines BSE and EDX analysis for identification and measurement of mineral concentrations, particle size distribution, and liberation/locking parameters. In addition, it can provide bright particle searches for PGM, gold-silver ores, and/or REE. With the newly developed Automated Loading System (AutoLoader), the TIMA has become the world?s first automated mineral analyzer which permits robust, continuous and unattended sample loading. With the AutoLoader (robotics) system, the TIMA has emerged as a powerful plant support instrument. It increases the sample throughput, minimizes manual labor and enables 24/7 operations. As such, it can be used in daily, continuous mineralogy for mine geology/ore control, concentrators, leach operations, and smelter support. It will also increase the capabilities of technical service centers for mining companies, commercial laboratories, large drilling programs or pilot test campaigns. In combination with a highly automated sample preparation lab (sizing, potting and polishing), TIMA technology can establish high-throughput and fast-turnaround automated mineralogy lab modules for mine-site production support.

Feb 23, 2014

By George B. Griswold

The two most stringent parameters for in place fracturing of rock are the lack of any initial void space in the rock which to expand into and the ability to gain access only through drill holes. These two restrictions have proven so severe that as of today the only proven techniques for efficient fracturing are the contained nuclear detonation and hydrofracing. Both of these methods have highly developed technologies for their application. Hydrofracing has been a standard oil field practice for two decades, and the experience gained from a couple hundred underground nuclear detonations has resulted in a thorough understanding of the phenomenology of cavity and chimney formation. What is still lacking is a practical method to create a suit- able degree of fracturing for those ore bodies which do not meet the application criteria for either the nuclear explosion or hydrofracing. Nuclear applications must be restricted to thick ore deposits so as to take advantage of chimney collapse height--preferable in excess of 200 feet thick. Other requirements such as ground shock, containment, and radioactive contamination further reduce the number of sites that prudently can be selected.

Jan 1, 1969

By John C. Stam

During the past 20 years, the seismic refraction method has been used for depth to bedrock investigations. Its usefulness, however, has been recognized mainly in engineering applications, for instance in connection with the construction of power stations and in dam site investigations etc. Although some knowledge of the depth to bedrock can be of advantage in many other fields, the use of the conventional 12 or 24 channel seismic refraction equipment has been very limited, especially in ore prospecting and mining operations. The reasons for this are mainly the rather complicated operation of the instruments and the high costs. A conventional seismic equipment requires a crew of six men, a highly trained operator and an interpreter. Furthermore, the use of explosives can under certain conditions be undesirable (naighbourhood of houses, damage to agricultural areas etc.). Since 1957 a new one-channel seismograph, 'constructed by Geophysical Specialties Company, Minneapolis, Minnesota, has been in use and this instrument simplifies operations and reduces their costs considerably.

Jan 1, 1959

By W. –E. Marx, M. Faber, R. Pollak, I. Astberg

INTRODUCTION Essentially three strategies are pursued in parallel for protecting the underground working force in mining against the consequences of methane ignitions which might lead to explosions: • Sufficient ventilation for avoiding the formation of explosive methane-air mixtures. • Avoiding sources of ignition caused possibly by electrical equipment or friction. • Structural explosion protection for limiting explosion propagation to the farthest-going extent for the case that methane is fired because of failing of the above preventive measures taken. The corresponding methods taken in mechanized roadheadings with point-attack roadheaders in German coalmining are subject of this paper. VENTILATION In ventilation of roadheading sites the following is to be considered in respect of methane control: • Sufficient air flows by volume being cycled to the heading face for diluting released methane to safe concentrations. • Sufficient local air velocities for effective mixing of methane and air in the whole heading site, in particular directly on the heading face and underneath the roof. Furthermore, ventilation needs to be laid out in a way that cut- ting does not result in health affecting dust exposure at any working place. Accordingly, the ventilation systems shown on Fig. 1 including primary and, if necessary, secondary auxiliary ventilation air ducts as well as dedusters, are used in roadheadings. This equipment is run according to the following principles: • If the roadheader is at standstill, the heading face is ventilated by the free air jet from the blowing duct closest to the face. In this operation phase the deduster is shut-down. • During cutting, the exhaust of the blowing air duct is automatically closed, and the deduster is taken into operation. The in- coming fresh air is now locally discharged via a lateral orifice, or distributed by Coanda-type ducts from the primary or secondary ventilation air duct. In this way, a stable dust wall not disturbed by the jet of the air duct is formed in front of the heading face. Only the ventilation air stream taken in by suction of the close-to-face deduster is passed through the inby zone of the heading face (over approx. 20 m). The clean exhaust of the deduster is dischayed to the roadway in some distance outby. Immediately at the heading face, methane accumulations possibly released by cutting are diluted by compressed-air jets from air-movers located near the roof which assure more rapid motion of the air without distuhing the dust wall. • In the zone of overlapping of primary and secondary duct and the deduster - that zone possibly is poorly ventilated during cutting - sufficient local air speed is assured by Coanda-type ducts or additional ventilation jets for avoiding methane accumulation. The regulations of the inspectorate specifies geometrical conditions to be met with these configurations of ventilation equipment. In-seam roadways are ventilated at rates > 0,5 m/s. For dedusters, present-day ventilation air flows are of 600-800 m+/ min. Ventilation monitoring and control is assured by units for measuring methane and ventilation air flows. If pre-set thresholds are exceeded or fallen short of, these measuring units shut-down - dependent on potential hazard - any not intrinsically safe electrical equipment in the roadheading, either totally or in limited zones. In addition, CO-measuring units near the roadway en- trance serve for early detection of mine fires. In spite of this ventilation layout and the equipment used and in spite of the water spray systems on the cutting head (discussed later in this paper) several cases of methane ignition and subsequent burning off (however, no explosions) have been recorded in the recent past. Therefore, the above-described ventilation is nowadays complemented in many cases by machine-mounted ventilation equipment for the immediate surroundings of the cutting head. Ventilation of the surroundings of the cutting head Investigations carried out subsequent to the above-mentioned gas ignitions have shown that the high ventilation air speed necessary for rapid dilution of methane could not permanently be arrived at in the following zones: • The direct zone of cutting, in particular during sumping. • The floor zone above the debris, predominantly below the cut- ting head. Accordingly, VOEST ALPINE-Bergtechnik, a producer of mining machinery, in cooperation with DMT developed for point- alttack roadheaders a novel machine-mounted ventilation equipment, the Jet-Block" (AERO SAFE Jet-Block) shown on Fig. 2. This system allows to ventilate the above-mentioned critical zones with high flow speeds. One AS Jet-Block is mounted on each side of the boom approx. 0.5 m behind the cutting head. The blocks discharge compressed-air jets with high energy from 10 nozzles per block into the hazardous zones (Fig. 3). Two nozzles of each jet block are tangentially oriented toward the upper circumference of the cutting head, the others are arranged fan-like with orientation to the heading face underneath the cutting head and the debris. The nozzle array of the blocks as well as the com-

Jan 1, 1997

In 1981, North American Degerstrom Contracting began gold mining on the Little Rocky Mountain Range, near Zortman, MT. In the past four years, the price of gold has fluctuated from a high of $16/g ($500 per oz) to a low of $9.60/g ($300 per oz). With that kind of a moving target, profitability presents a management challenge. Based in Spokane, WA, Degerstrom employs 85 people at the mine. It is one of the largest heap leach operations in the nation. Last year, 8 Mt (9 million st) of low-grade ore was moved to the leach pads in a two-shift, five-day operation. The current pit being worked produces 218 kt/m (240,000 stpm) of leaching ore. Ore fragmentation critical The gold bearing ore has no defined patterns or seams. It is an igneous intrusion resulting from volcanic activity. It occurs spontaneously throughout the range. To recover the gold, areas are staked out and core samples are taken and analyzed. These data determine if the ore is rich enough to mine. Once mining begins, an average of 150 holes are drilled and blasted on a two-day rotation. Five Ingersoll-Rand T4s drill on a 4- x 4.5-m (14- x 15-ft) spacing to a 7.6-m (25-ft) depth. "When we first came to this site, we drilled on a 3.6- x 3.6-m (12- x 12-ft) pattern," said Paul Baker, superintendent for N.A. Degerstrom Contracting. "We did some experimenting using a high density emulsion in the bottom of the hole and filling the rest with regular Anfo." The high density emulsion saved the company on overall drilling and blasting costs since the holes could be spaced on a wider pattern and still achieve the high degree of fragmentation needed for leaching. Blasting is important because the gold lays in the natural strata of the rock. Also, the ore is not crushed before it goes to the leach pads. So complete fragmentation is critical. Terrain dictates loading/hauling system The shot ore is worked in 6 m (20 ft) benches by two Caterpillar 245 front shovels equipped with 3 m3 (4 cu yd) buckets, a 992C wheel loader with a 12.6-m3 (16.5-cu yd) bucket, and a 988E high lift fed by a D9H. A Caterpillar D10 tractor trap dozes to the loader. According to Baker, the 6 m (20 ft) bench is an efficient lift for the front shovel and the wheel loader. In tight quarters or in a pocket of ore, Degerstrom uses the 245s. For high production areas, the wheel loader is used. The hauling fleet includes 28 Cat 773 off highway trucks. The 45 t (50 st) trucks are seven pass loaded by the 245s with a 25-second cycle time per bucket load. The 992C loads the trucks in two passes in less than 30 seconds. The cycle time is four to five minutes for a 1.2-km (4000-ft) haul on grades that average 10%. Grades, in some places, exceed 16%. "We use the 773s because they have a low weight-to-horsepower ratio - that's what you need in steep country," Baker said. The trucks haul the shot ore to the leach pad. Currently, Degerstrom is working a pad that has a capacity for 5 Mt (5.5 million st) of ore. Building the leach pads The base of the pads resemble a large drainage basin. They take about one month to construct. A 0.3-m (1-ft) layer of impervious bentonite clay is hauled in and leveled by a Cat D9H tractor. A 30-mm (1.2-in.) PVC liner is then laid in place on top of the clay base. It, in turn, is covered with tailings from an old on-site mill. Degerstrom uses the tailings to protect the liner from tears when the ore is dumped. The ore is leveled and built up in 9 m (30 ft) lifts by the D9H. The tractor rips the top layer. Then, a network of plastic irrigation pipe is put in place to distribute a cyanide solution over the surface. The leaching solution percolates through the ore and dissolves the gold. The solution drains from the pads and is pumped to a 22.7-ML (6-million gal) pregnant solution holding pond. The liquid then goes through two separate filtration units. One unit removes entrained solids and the other side adds a zinc dust to precipitate the gold. This is collected on filters. The affluent then returns to a solution holding pond for redistribution through the pipeline network. Typically, 80% of the pad's potential recovery takes place in the first 30 days of leaching. The process stops when the ambient temperature falls below the cyanide's freezing level. And each pad has a leach cycle of four to five years before recovery values decline to a point where further leaching becomes uneconomical. At Zortman, it takes 9 t (10 st) of ore to produce 31 g (1 oz) of gold. It is therefore critical to keep total production costs down and efficiency high.

Jan 1, 1986

By S. L. Bessinger

Introduction In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway's complex geometric cross section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Head losses in mine entries are calculated using Atkinson's Equation. [22H= KPLQ (English) H= KP 3Q (SO (1) 5.2AA] where H = pressure loss, in. of H2O (Pa); K = friction factor, lbf•min2/ft4 (kg/m3); P = perimeter, ft (m); L = airway length, ft (m); Q = airflow quantity, ft3/min (m3/sec); and A = flow cross-sectional area, ft2 (m2) In this equation, the friction factor, K, is an empirical constant that describes the aerodynamic roughness of the airway. Typically, the K-factor for a given airway is determined by measuring the factors H, P, L, Q and A in Equation (1) and calculating K. Tables of friction factors calculated in this way are found in textbooks and handbooks that deal with mine ventilation analysis. Unfortunately, very few K-factors have been measured on longwall faces, and the accuracies of those that have been measured are entirely site specific, because of the wide variety of equipment found on longwalls. The development of a technique for prediction without mine-site measurements of the friction factor for any longwall face, regardless of its configuration, will thus be very useful in the design of ventilation systems for mines in which longwall mining is practiced. Calculation of pressure losses using Atkinson's Equation (1) and empirically determined K-factors provides accurate and useful approximations in cases where the airways have relatively simple cross sections. However, a careful analysis using the principles of fluid mechanics shows that such calculations are based on two assumptions that are not strictly correct when there are obstructions in the airway. The first assumption is that the air velocity distribution in the cross section, particularly around the perimeter, is uniform. This assumption results from the fact that the tabulated K-factor values found in the literature are based on field measurements with uniform conditions. Such uniformity does not exist in longwall airflows. The second common assumption is that the K-factor, and corresponding head loss, is independent of the Reynolds Number (NR) for a given flow. In fact, this assumption is not strictly correct, and is particularly erroneous in the case of irregular protuberances into the airflow, such as those found on a longwall face. The errors arising from the assumptions may be avoided by using K-factors calculated by a newly devised method, described below, which takes into account the fundamental principles of aerodynamic drag analysis. This new technique has two advantages: first, it is flexible enough to model any longwall, regardless of equipment configuration; second, it employs terminology and equations familiar to those who perform mine ventilation analysis, using K-factors, for which ventilation engineers have an intuitive understanding, rather than drag coefficients. To provide guidance for development of a longwall drag model, data were taken on two modern longwalls operating in substantially different conditions. Pressure measurements at Mine B were made with 200-foot (61-m) sections of 1/8-in. (3-mm) diameter plastic tubing, attached to a Dwyer Magnehelic gauge. Pressure drops were measured in 200-foot increments down the face, and summed to give the drop for the entire face length. This technique was found to produce small, repeated errors because of the number of segments required to span the longwall. At Mine A this problem was avoided by using a single, continuous, plastic tube for the entire face length. The psychrometric properties of the air were measured for both Mines A and B. A calibrated, standard vane-anemometer was used to measure the airflow on both faces. Finally, numerous dimensions were measured on both faces, and face profile drawings were obtained to allow detailed evaluation of the face equipment geometry. From this information, accurate evaluations of the average wetted perimeter and average area of the longwall face airways were made. Since the airflow is not confined to inside the powered supports at all points along the face, a quadratically weighted average of the airflows measured at various stations along the face was calculated: [n2Qavg =Qi Ii / It(2)i=1] where [Q, avg = average airflow for analytical purposes, ft3/min (m3/sec); Q= airflow at station i, ft3/min (m3/sec); 1= length of segment represented by Q, ft (m); h= length of longwall face, ft (m): and n= number of quantity measurement stations.] The quadratic weighting scheme was chosen because the

Jan 1, 1992

By J. E. Oberholtzer, M. G. Fernald

INTRODUCTION Approximately six years ago the U.S. Bureau of Mines (USBM) developed a data acquisition system (DAS) specifically designed for measuring radon levels and other environmental parameters during studies of means to control radiation hazards in underground uranium mines. The DAS system records data in machine readable form using a paper tape punch, which represented the state-of-the-art at that time for a moderate cost output device. However, the use of paper tape as a recording medium for field studies is somewhat unwieldy. Reducing the raw data required either that the tape be shipped to a computer center equipped with a high-speed paper tape reader or that the tape be transmitted at low speed over the telephone lines to a remote computer. Transmitting, at ten characters per second, the data from a 10-channel DAS taking Four readings per hour would require about 30 minutes For each 24-hour day's data. Telephone lines from remote mine sites are often of marginal quality and data errors can be introduced during transmission. Paper tape punches are also prone to occasional punching errors. Both problems make it necessary to carefully check for and correct data errors, a process which is possible because each DAS produces an independent printed data record, but the error checking and correction process can be quite laborious. Aware of recent advances in microcomputer technology which have brought the price of a personal computer down to about the cost of a paper tape punch 5-10 years ago, the Bureau decided to explore the feasibility of using a low-cost personal computer in the field to process DAS data in real time. On behalf of the Bureau, Arthur D. Little, Inc., developed a simple interface circuit which permits an Apple II computer to accept data from one or two DAS units as it is being transmitted to the paper tape punches. Computer software converts each measurement to appropriate engineering units, e.g., radon concentration, Working Levels, air velocity, temperature, or barometric pressure. The computer also calculates 1-hour and 8-hour running averages of all converted data and prints those results as soon as they are obtained on a line printer located at the test site for immediate inspection. After development, the system was used continuously and successfully for a 5-month period at a Utah uranium mine. DAS DESIGN AND MODIFICATION Each of the two USBM data acquisition systems used in this work consists of two separate modules. A multiplexer module located below ground near the measurement transducers acquires signals from each of nine tranducers. Six input channels were devoted to measurements of radon or Working Level. The outputs of those transducers, photomultiplier tubes or G-M tubes, respectively, are digital pulse trains which are accepted directly by the mutliplexer. Three channels were used for environmental parameters--air velocity, temperature, and/or barometric pressure. Each of the environmental tranducers is fitted with dedicated linearizing and voltage-to-frequency conversion circuitry so that the outputs to the multiplexer are also pulse trains having frequencies of one tenth of the value of the measured parameter expressed in the appropriate engineering units. A 100-Hz reference signal was input into the tenth channel for use in monitoring system integrity and performance. All ten pulse trains are then timeseries multiplexed into a signal line for transmission to the above-ground data acquisition module. Above ground, the composite signal is de-multiplexed into ten separate lines, each of which is connected to a digital counter which converts the pulse train to a numerical value. The acquisition of each set of readings is initiated by an adjustable "scan cycle comparator" timer. The acquisition process proceeds in three phases. First, radon and Working Level channels are counted for an extended period of time, typically 5-10 minutes depending on activity, because of the low pulse rates involved. Then the other four channels are counted for ten seconds, and finally, all ten readings, along with the Julian day and time of day are output serially onto paper tape and printed on a strip printer. When the scan cycle comparator reaches its preset time (15-minute cycle times were used in this work), it resets itself, initiates another readout cycle, and begins timing again. The only modification made to the data acquisition systems used in this work was to disconnect the scan cycle comparator in one unit, which became the "slave" and bring in the scan cycle comparator signal from the other unit, the "master", to initiate data acquisition cycles in the slave. Synchronizing the two data acquisitions in this fashion and using two slightly different radon counting times insured that the two systems never attempted to output data to the Apple II at the same time. THE APPLE II COMPUTER The Apple II computer used in this work was equipped with 48 KBytes of semiconductor random access memory (RAM), two floppy diskette drives, a Centronics Model 730 impact matrix printer and a modulator for driving an ordinary color television as a video display device. A single California Computer Systems Model 7720A dual 8-bit bidirectional parallel input/output (I/O) card was installed in the Apple to accept the digital data from both data acquisition systems. This card is

Jan 1, 1981

By G. L. Schroeder, C. H. Summers, D. B. Lindsay

INTRODUCTION The requirement that is placed on owners and operators of underground mines to protect workers against the health hazard of inhaling radioactive materials which are short-lived decay products of 222Rn can be satisfied by applying a considerable variety of what we may call "engineering" solutions as well as a number of "administrative" remedies to the problem. The most obvious of the "engineering" approaches has always been that of forced ventilation, in which relatively clean (i.e., radon-free) air from aboveground is drawn or pushed through the mine workings by a system of fans and ducts. This relatively clean air, in sweeping through the drifts, stopes and haulageways, dilutes the radon and radon-daughter concentrations in the air of the mine, and performs the added beneficial function of removing the daughter-mixture quickly enought to limit grow-in of the longer-lived nuclides in the group that make up the "toxic trio" on which the Working Level (IM) unit is based. Effective as the dilution-ventilation method is for control of WL in most underground mining situations, however, the increasing strictness of control measures that have been imposed on the mining industry over the last two decades have demanded measures of even greater effectiveness. In times of poor markets for yellow-cake and other products of the mines, mine operators are pressed to reduce operating costs, and the installation of additional primary ventilation capacity can be a severe burden on a mine that is already laboring under an unfavorable earning power. When traditional dilution-ventilation systems alone cannot meet the requirement for WI, control, radiation safety engineers and ventilation engineers begin to look at alternatives and auxiliary methods. Since the radon which produces the toxic daughter products originates in the rock of the mine walls, and since, in most United States mines, that rock is a porous sandstone through which air can move under the effect of atmospheric pressure gradients, and through which radon can diffuse relatively freely, one way to help control the growth of WL would be to hinder the escape of radon from that reservior of porous rock. An appealing; method for hindering that natural flux of radon-polluted air from the walls of the mine has long been apparent; namely, to apply a low-permeability coating over the surface of the rock, thus sealing the radon in place and, in theory at least, preventing its escape into the mine air. Our 1970 report to the U.S. Federal Radiation Council on the subject of cost impacts of proposed changes in the occupational standards for exposure of underground uranium miners to airborne radon daughters noted the possibility of using polymeric wall sealants as a means of controlling radon-pollution of mine air. Since that time a number of reports have appeared in the technical literature advocating this technique for restraining the escape of radon from building materials, mill tailings, and other materials containing 226Ra, in addition to the surfaces of underground mine workings. During this period, some controversy has occurred over the question of the probable effectiveness of wall sealants in limiting the escape of radon from the rock. Our 1970 report speculated that flaws (cracks and "pinholes") in the coating might be all but unavoidable in practice, and that even a conservative estimate of the frequency of such flaws would lead to a prediction of ineffectiveness. Hammon et al, in a laboratory evaluation of radon sealants conducted by Lawrence Livermore Laboratory of the University of California in 1975 on behalf of U.S. Bureau of Mines, concluded that a wide variety of polymeric coatings would provide "nearly 100% effectiveness" in restrain¬ing escape of radon from mine wall surfaces if applied in "thicknesses between 5 and 10 mil" (125-250 [y]pm). John Franklin and co-workers at the U.S. Bureau of Mines laboratories in Spokane, Washington, have carried the experiments with polymeric sealants through additional laboratory tests and into actual mine environments, reporting that selected sealants could provide attenuation of radon flux by a factor of four (75-80% reduction). Robert Bates and John Edwards of USBM developed a computer-assisted mathematical/physical model that predicts a relatively small effect of flaws in a low-permeability coating on the radon flux from a sandstone-type matrix. FIELD TEST Since all actual experimental work with wall sealants showed some beneficial effect on radon attenuation (even if not as exciting as the "nearly 100%" predicted by Hammon), USBM was encouraged to extend its evaluation to an actual operating uranium mine, and awarded a contract for that work to Arthur D. Little, Inc. in September 1979. We were fortunate in obtaining the voluntary cooperation of Atlas Minerals Division of Atlas Corp., who operate a mill and several underground mines in and around Moab, Utah. Atlas made available for our use a small T-shaped drift in their Pandora Mine in LaSal, Utah, and provided space for instrumentation and recordkeeping by our field crew in a surface building near the mine entry. Atlas also provided electricity and water to the test site, together with assistance in establishing necessary ventilation, removing rubble from the site, conducting periodic WL surveys and furnishing auxiliary man-power for the heavy hard work of coating the walls with gunite prior to application of the polymeric sealant. The generous coopera-

Jan 1, 1981

By Ralph P. Hackl

Ore at the Congress gold property, located near Goldbridge, British Columbia, is extremely refractory with most of the gold disseminated in fine-grained arsenopyrite. Metallurgical testwork conducted by Giant Bay Biotech Inc. indicated that biooxidation pretreatment was capable of increasing cyanide gold recovery from the ore to greater than 90%. To determine if the bioleach process could be successfully scaled up, a 225m' (7950-cu ft) commercial size bioleach reactor was built and tested at the mine site in 1989. At that time, the tank was the largest bioleach reactor ever operated. The objectives of the program were: • to ascertain if the biooxidation rates obtained in lab and pilot-scale reactors could be duplicated in a commercial scale tank; • to identify agitation conditions under which the process will function optimally; and • to demonstrate the bioleach process on a scale sufficient to satisfy the owners and potential investors.

Jan 1, 1990

By Patrick M. Hurley

The spreading-ridge, or accreting plate margin, is the primary conduit of metals delivered to the earth's surface. Isotope ratios of lead show primitive mantle values at these sites, indicating that this is true. We can use lead isotopes in any mineral deposit to indicate the degree to which the metal is "primitive" in the sense of having been uncontaminated with U-enriched crust or "crustal" in the sense that it has been cycling at the earth's surface since earlier times. The reason for this is that the upper continental crust has an enrichment in the ratio of U/Pb over that in the mantle, so that old continental leads have a larger radiogenic component. The primitive metals delivered to the surface at the spreading-ridge are transported on the oceanic plate until they reach a subduction zone. Here they become cycled in the "orogen". However, in this process the primitive metals are mixed with metals from the continent so that the average lead isotope ratios are between those in the mantle and the average upper continental crust. Subsequently a part of the metals in the orogen may become fixed in a new block of continent. However there are processes which continue to uplift and erode ancient continental regions so that there is a continuous feeding back of metals to the orogen. We have therefore a cycling of metals from the existing continental crust into the orogen where they are mixed with new contributions from the spreading ridge and mantle.

Jan 1, 1977

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

The STG integrated method of grouting can be divided into the several phases as outlined in preceding chapters. For convenience in describing technical procedures, they can be grouped as: 1. Investigations in non-geotechnical, exploration bore¬holes. 2. Drilling the holes necessary for geotechnical investiga¬tions and for grouting. 3. Conducting appropriate hydrogeologic tests in those holes and consequent calculations. 4. The preparation of a clay mortar. 5. The preparation of a clay-cement grout with additives. 6. The injection of the grout. 7. Checking the quality of the grout curtain. Items 2, 4, 5 and 6 are discussed below. These activities contain the procedural aspects of grouting. 7.1 DRILLING GROUT HOLES Grout holes belong to a very select class of drillholes. The necessary drilling equipment must enable the operator to drill the holes in inclined, twisting directions. In addi¬tion, the equipment must permit drilling under structurally and hydrogeologically complex conditions. The equipment must permit the operator to conduct specific testing activi¬ties in the drillholes. These activities include hydrodynamic analyses, flowmetric analysis, and the installation and re¬moval of deflectors, packers, grouting plugs, casings, lin¬ers, pumping facilities and other work. As explained in Chapters 5 and 6, the grouting of sat¬urated rock is conducted both through holes drilled from the surface and through holes drilled from the face of a shaft, drift, or tunnel. The drilling of grout holes from the surface can be carried out by an aggregate of equipment that can be the same equipment used for drilling exploration boreholes. Depending on the projected depth of the holes, the STG ZIF- 1200MR and ZIF-650 drilling rigs (or modifications of them) are used. The SKB-4, SKB-5 and SKB-7 high-output drilling rigs (workover rigs) have been used in recent years. The STG BMP-24 drilling rigs are used for carrying out the descent-lifting operations for drilling holes from the sur¬face. In a number of cases, the UKB-500C, URB-3AM and URB-2A power-fed drilling equipment is used for drilling holes from the surface. It should be noted that the drilling of grout holes usually requires an electric drive assembly. Drilling with diesel drive assemblies would be used only to preclude the possi¬bility of losing electric power during drilling operations. Turbine drilling would be used only for appropriate techni¬cal-economic reasons. The selection of the grout hole design is determined by the hydrogeological and structural geological conditions at the site as interpreted in Chapters 2 and 3. The design of grout holes must be as simple as possible in order to min¬imize costs. Hole design guidelines can be adopted as a basis for this purpose, unless site-specific conditions require otherwise. The mouth of the drillhole must be outfitted with a guide-pipe having a length of at least 2 m and an outside diameter from 219 to 234 mm. The upper part of the hole must be attached by a jig consisting of borehole casing that has a diameter ranging from 108 to 146 mm. The length of the jig h is determined by the equation [ ] where k = 1.1 to 1.2 is the load factor; P,, is the injection pressure of the grout into the mouth of the drillhole; D is the external diameter of the jig pipe; m = 0.6 to 0.7 is the work condition factor; T[ ].1 MPa is the bonding value of the rock cemented to the jig. The grout hole is drilled from the jig shoe to the de¬signed depth using a rock-crushing bit with a diameter of 93 mm or more. However in complex hydrogeological condi¬tions when the shaft intersects unstable rock, other hole designs can be used. The diameter of grout holes must permit an aggregate of investigations to be conducted in them using down-hole borehole geophysical and flowmetric logging instruments. It is necessary to cement the casing strings reliably in the grout holes, thereby permitting the injection of the grout under high pressure through the pres¬sure tight mouth of the borehole. Technological details for drilling grout holes such as the rotational rate of the drill bit, the axial load on the rock¬crushing bit, and the drilling fluid flow rate are optimized for the specific rock-hydrogeological conditions largely by experience. The hole diameter and type of rock crushing bit are important variables in these details of drilling. It is advisable to drill holes using water as the drilling fluid. However in those cases where the hole walls are unstable or under very high ground water pressure, the use of a well¬ engineered drilling mud should not be precluded. Unfortu¬nately, drilling mud can influence test results. When steeply dipping fractures with inclination angles of more than 60 to 70° are encountered, it is necessary to drill guided, inclined grout holes. This procedure permits the maximum number of fractures in each hole to be stu-

Jan 1, 1993

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

6.1 PRINCIPLES OF PLANNING GROUTING OPERATIONS FOR SHAFTS, DRIFTS AND TUNNELS The planning of integrated grouting of vertical, inclined or horizontal underground workings in saturated, fractured rock includes the following phases: 1. Planning for the collection and analysis of geological information concerning the characteristics of the struc¬tural deformation(s) that produced the details of the frac¬ture system being studied. 2. Planning for the acquisition of initial hydrogeological information produced by the flowmetric and hydrody¬namic investigations (described in Chapters 2 and 3). 3. Computing the initial engineering calculations of the ex¬pected or planned dimensions of the isolation curtains for each water-producing stratum (fracture zone) and the expected areal spreading pattern of the grout in each of these water-producing horizons. The selection of the lo¬cations, the initial optimal number of holes and the vol¬ume of grout needed to create each isolation curtain also are included in this phase. 4. The selection or development of a grout formula for the specific hydrogeological conditions that exist at the site. 5. Identification of a resource (deposit) required to produce the grouting clay. The planning of grouting operations must be supported by the acquisition of all necessary initial data, which in¬cludes: 1. The geographic location and administrative responsibil¬ity (ownership and regulatory constraints) of the mineral deposit and any peculiar features of the area where ex¬cavation of the underground workings is planned. The surface relief, the hydrographic network of the region, and the local climatic conditions must be evaluated prior to planning for buildings and waste discharges if any. 2. The economic status of the region, the available trans¬portation, the nature of the energy base, the water and heat supply sources, the location of construction mate¬rials suitable for making the grout all must be deter¬mined. 3. The geological and hydrogeological characterization of the work site must be accomplished. 4. The technical features of the shaft(s) and underground works must be characterized. 5. The organization's principal initial design plan features and the construction sequence of all buildings and struc¬tures must be developed. These must be correlated with the organization and schedule for excavating the under¬ground works. The technical characterization of a shaft or of other un¬derground workings must contain information about the purpose of the workings, the design depth, the planned excavation procedures along with cross-sections, and the parameters of the lining, if a lining is required. The geology of the section to be penetrated must include geological data obtained from existing exploration activities at the mineral deposit and conclusions about the shaft ex¬cavation conditions obtained from drilling technical moni¬toring holes and exploration drillholes. The interpreted stratigraphic regime, including the thickness, strike and dip of each stratum, must be shown in cross-section. The tec¬tonics, including all geologic structures, must be presented by cross-sections. The bedding orientations must be used to identify anticlines, synclines, monoclines and faults. The lithologic composition and the drillability of the rock com¬prising the section must be examined. The character of the rock's fracture pattern must be presented. Finally the per¬meability, porosity and storage characteristics of each aqui¬fer must be determined by in situ hydraulic property testing (injection tests or pumping tests, as described in Chapter 3). The fracture pattern of the indurated rock is presented in reports derived from geo-exploration borehole studies and from studies of outcrops on the surface and in all nearby underground workings. The extent of each fracture zone is mapped in three dimensions. The distribution of fracture apertures, the nature of the fractures according to each litho¬logic stratum, and the relationship of bedding planes to each fracture system are all determined. Acoustic logs, downhole cameras, caliper logs, oriented core, and any other appro¬priate geophysical tools available are used to construct the data base for this purpose. The hydrogeological characterization of the geologic section to be penetrated by the workings is based on the analysis of hydrogeological data collected during the drill¬ing and testing of technical monitoring holes and from in¬formation on ground water inflow collected during the ex¬cavation of nearby underground workings, if such workings exist. The hydrogeologic characterization must contain the following information: 1. The identification of all aquifers, their thicknesses and their locations in the section, 2. The distribution of hydrostatic head in the ground water in three dimensions, 3. The coefficient of saturated hydraulic conductivity (per¬ meability) of each aquifer, 4. The porosity and storativity of each aquifer, and 5. The projected rate of water inflow into the shaft, drift or tunnel for each aquifer prior to grouting. This step is used to identify the zones to be grouted.

Jan 1, 1993

By J. Gauntt

Calcium carbonate is one of the most widely used industrial minerals in applications like cement, lime, agriculture, power plant fluegas desulfurization (FGD) and construction aggregates. Most applications are generally high-volume and low value ($9/t or $10/st) or less free-on-board (f.o.b.) mine site) uses of this material. Specialty uses of ground calcium carbonate (GCC) are typically referred to as ?performance mineral? or ?extender and ?ller? applications. Primary specialty markets for GCC include paper and plastics as well as a number of construction applications including joint compounds, carpet backing, paint, and adhesives, caulks and sealants (ACS). Other important specialty applications for GCC include rubber, food and personal care products. A new and growing application for GCC is in nonwoven fabrics, where this mineral replaces a small percentage of the polymer. Typical specialty applications for GCC are color sensitive and require extremely white/bright materials. Although deposits of limestone are located in most states, high brightness resources are relatively rare and controlled by a limited group of suppliers.

By D. R. Gagnon

Mining companies are continually looking for product-ivity and efficiency improvements with their equipment to reduce operating costs. This paper will review the results of a productivity analysis conducted at the Luscar Ltd. Boundary Dam and Bienfait mines near Estevan, Saskatchewan using a Computer-Aided Earthmoving System (CAES) installed on a track-type dozer. The CAES TTT system is a Global Positioning System that provides the operator with constant survey guidance. The study includes a productivity comparison between two similar test sites, one with the aid of the CAES system and the other, without. The CAES TTT system shows a productivity improvement of between 20 and 35% in the reclamation slope leveling application. Even with the CAES TTT system deactivated, some of the advanced features of the system were used to: retain terrain data for volume calculations; track the equipment position; and, accurately record and separate idle time from working time.

Jan 1, 2002

By Paul B. Queneau

Formulation and commercial utilization of solid-waste incinerator slag by-product is the focus of this paper. Particular attention is given to iron aluminosilicate melts having compositions similar to that of nature's basalt The earth?s crust can be viewed as an environmental laboratory for predicting stability of various silicates; small-scale laboratory experiments are necessary to confirm findings there from. An ideal slag consumes flux additives at acceptable cost, is compatible with furnace refractory, flows readily from the furnace, and has physical properties suitable for its commercial utilization -all within the critical constraint that the slag is environmentally stable. Examples of successful solid waste slagging practice, including tonnage, utilization and pricing, are provided from metallurgical and waste treatment operations of North America. Western Europe and Japan.

Jan 1, 1991

By A. Adu-Acheampong

The Mid-Tennessee Zinc (MTZ) mines, comprised of the Gordonsville, Elmwood, and Cumberland sites, last operated 4 years ago and are undergoing rehabilitation to begin production in 2008. The mines utilize a random room-and-pillar mining method that results in complex ventilation networks. Upon decommissioning, the mechanical ventilation units were removed and disposed of and several shafts sealed. In an effort to re-open the mines and operate in compliance with imminent Mine Safety Health and Administration (MSHA) diesel particulate matter (DPM) regulatory limits of 160TC µg/m3, a new ventilation system was required. The new ventilation system was developed primarily based on the diesel equipment required to meet production. Particular emphasis was placed on the DPM, diesel equipment utilization, main fan sizing, and fan locations to enable the mines to meet MSHA?s ventilation requirements in a cost efficient manner.

Jan 1, 2009

By T. H. Chan

As the University of Hong Kong and their advisors planned their new Centennial Campus it became clear that they would have to re-provide existing water reservoirs on an adjacent site with new salt water reservoirs located underground in two new caverns. The existing reservoirs are owned by the Water Supplies Department who provide salt water for sanitary flushing and a separate potable water supply for regular use. Although underground reservoirs had not previously been adopted locally, the planning has been completed to a tight schedule and the construction is well advanced. The need to preserve historic buildings imposed constraints, as did the requirement to minimise potential construction nuisance for the local community and environment. This paper presents the background to the project and the work carried out to develop an innovative, sustainable and environmental acceptable solution. Construction progress to date is also presented.

By Richard J. Ebens

The trace element geochemistry of the cherty, clay-rich carbonate residuum. Covering much of southern Missouri and adjacent parts of Arkansas was investigated during the early 1970's as part of a statewide geochemical survey of Missouri-1/2/3 The study was environmentally oriented and focused on the character of the "natural' geochemical environment of the state. A great deal of effort was made to collect samples that would represent the geochemical variability of the near-surface environment in an unbiased way. In general, this was accomplished by use of randomization procedures in both collection and analysis. Most of the sampling sites were selected randomly to prevent personal bias from entering into sample selection. In the laboratory, samples were analyzed in a randomized sequence to circumvent the effects of any systematic bias that may have entered into the laboratory procedures. Several residuum samples contained 1 000 parts per million (ppm) or more of Pb, Zn and Ba. All these samples were collected in Washington County in or near the barite district. The frequency distributions shown in figure 1 indicate just how unusual the residuum in Washington County is with regard to these three metals and copper. The shaded parts of the histograms represent the samples from Washington County. The histograms are based on analyses of 141 samples from 38 localities scattered over the subcrop of the Cambrian and Ordovician Formations of southern Missouri and northern Arkansas. Eight localities were selected randomly from over each of four bedrock units--Bonneterre Dolomite, Potosi and Eminence Dolomites, Roubidoux Formation, and the Jefferson City, Powell and Cotter Dolomites. Six localities were selected randomly from over the Gasconade Dolomite. The geochemical "background" for these elements in residuum is defined by samples from outside of Washington County (unshaded part of the histogram); lognormal distribution curves have been fitted to the corresponding parts of each histogram. Clearly, these lognormal models (curves) describe well the background

Jan 1, 1977