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By J. Guzman

Introduction The Carlin Trend is North America's premier gold producing district. It is located in northeastern Nevada's Elko and Eureka Counties along a northwest trending belt about 65 km (40 miles) long and 8 km (5 miles) wide (Thorstad, 1989; Jones, 1989). This trend is the worldwide reference site for epithermal, sedimentary rock-hosted microscopic gold deposits. At least 19 deposits have been discovered to date, varying in size from 933 kg to 1.08 kt (30,000 to 35 million contained oz) of gold (Fig. 1). Newmont Gold Co. and its parent, Newmont Mining Corp., jointly constitute the largest mineral right holders in the district. They own or control more than 1000 km' (386 sq miles) in and around the Carlin Trend and own all or part of I6out of the 19 mines and prospects identified to date. Since the initiation of Newmont's exploration activities in the Carlin Trend in 1961, 2.24 kt (72 million oz) of cumulative gold resources have been identified. Cumulative production from all mines since the start-up of Newmont's Carlin mine in 1965 to the end of 1989 was about 202 t (6.5 million oz) (Jones, 1989). The incentive of sustained high gold prices and innovation in processing technology resulted in a significant acceleration of gold output over the last few years. Newmont Gold alone produced more than 43.5 t (1.4 million oz) in 1989. That is equal to 22% of the cumulative 1965 to 1988 output, and an almost 200% increase over its 1986 output. The same incentives produced even more spectacular exploration results. In each of the last five years, net additions to reserves and resources outpaced current production by substantial margins. These facts demonstrate the spectacular past prospectiveness of the Carlin Trend and the success of focused, multi-disciplinary exploration methods that made it possible to more than offset the recent accelerated depletion of gold resources. However, is this situation sustainable? How long can the mining companies along the Carlin Trend keep on finding resources faster than they deplete them? These are some of the questions that motivated this study. The authors have not quantified the future potential for gold exploration in the Carlin Trend nor established a deposit discovery path. But strong indications were discovered that the [ ] Carlin Trend remains a relatively immature exploration district and that the potential for significant new discoveries is high. Methodology and data The approach chosen to address the above questions was simple. The authors compiled deposit size data, measured in contained ounces of gold resources, for all known deposits along the Carlin Trend (Table 1). The resource information was obtained by adding cumulative historical production (adjusted for mining losses and metallurgical recovery) to 1989 year-end published resource inventories. In a mature exploration area, where most deposits have been discovered, this distribution would be expected to approximate lognormality and would plot along a straight line on a lognormal probability scale. This result was found in previous work by Allais (1957) and recently confirmed by Cox and Singer (USGS, 1986) in regard to various types of mineral deposits in several regions of the world. It was also found to hold true for oil and gas pool size distributions (Arps and Roberts, 1958; Kaufman, 1962; McCrossan, 1969). [ ] The data used were compiled by Newmont Exploration geologists. The purpose of the study is to make inferences about the underlying geologic processes in the district and the maturity of the exploration effort. Therefore, deposits were not classified according to ownership but according to geologic occurrence as known from current information. Newmont's Post and Barrick's Goldstrike and Betze deposits, for example, are shown as a single occurrence to reflect the actual geologic setting. The cumulative frequency distribution of deposit sizes was plotted on lognormal probability paper (Fig. 2). The abscissa shows the cumulative fraction of deposits at or below a certain deposit size and the ordinate shows the deposit size in thousands of ounces of contained gold resources.

Jan 1, 1992

By Don Suttie, Frank Roybal, John Wright

The Roybal Raise incorporates the simplicity and low cost of a short bald-headed raise while overcoming the usual impediments to driving a high baldheaded raise. The concept provides for speedy excavation and maximizing the use of the excavated opening, using the rock for structural support. The design is simple: two baldheaded raises are driven parallel to one another, interconnected by dogholes about 8 m (25 ft) apart. In this fashion, the raise can be driven safely and efficiently to 60 m (200 ft) or more. When the raise is complete, the excavation requires no timber for support or partitioning. Rather, the rock itself supports the construction. As a test, a two-compartment Roybal Raise was attempted at the Independence mine in the Willow Creek District of Alaska. One side was driven 2 x 2 m (6 x 6 ft), intended for a manway and equipment skip. The other was driven 1.5 x 1.5 m (5 x 5 ft) for a muck compartment. Both sides were driven at a 65° angle from horizontal and 1 x 1 m (4 x 4 ft) dogholes spanned the 4.6-m (15-ft) pillar between the two compartments. After topping out at 50 m (160 ft), the raise was found to meet all expectations of efficiency, low cost, and practical simplicity. Technique Preparing the Roybal Raise for driving involves no elaborate or costly site prep except assembling materials and providing for muck removal. Normally, the raise muck falls directly onto the sill or track and is picked up either by an overshot mucker or a front-end loader. Raise mining proceeds initially by taking up both compartments simultaneously to the first doghole. It progresses by alternating one side as a clear manway for access with the other side, the active side, as a muck compartment (Fig. 1A). This stage proceeds quickly for no mucking or staging is required to cycle the first three rounds. On the fifth round, in addition to the normal raise round, rib holes are winged out toward the adjacent raise, thus starting the dogholes. Finally, as the sixth raise rounds are drilled and shot, the first doghole is also completed from both sides. The second stage commences by establishing one side as a clear manway for access while the other side, used as a muck compartment, is prepared for active mining. Figure 1B illustrates this step in the succession of raising with the active side advancing to the second doghole level. During this stage, services are maintained through the manway side and equipment is stowed in the doghole during blasting. As the active side advances, muck may be allowed to fill the bottom portion of that side since it is unnecessary to clear it for access. The next stage of raising involves cleaning out the former muck compartment and preparing it for use as a manway. At the same time, the former manway is prepared for advance and use as a muck compartment (Fig. 1C). As in the previous step, the same doghole is used for access and equipment storage. The active side advances until the second doghole is completed, at which point equipment and access are moved to the new doghole. From there, the active side is driven past the second doghole 8 m (25 ft) more to the third doghole level. Again, half of the doghole and the next raise round are mined. The fourth stage involves switch ing back the manway side for the muck side (Fig. ID). The new muck side is driven to the third doghole where equipment storage and access are reestablished. From there, the active side is advanced to the fourth doghole. In the same fashion, the raise is advanced to the top, alternating the manway side with the active side, maintaining access through the inactive manway, and storing equipment in the uppermost doghole. The cycle time for driving each compartment, regardless of the raise height, is greatly improved over that for normal bald-headed raises of comparable height. By storing equipment in the upper-

Jan 1, 1984

By R. H. Moore, B. D. Breitenstein

INTRODUCTION The United States Uranium Registry (USUR) tissue program was funded in 1978 by the United States Department of Energy (DOE) through its Human Health and Assessments Division, and registrant enrollment began in 1980. It is operated by Hanford Environmental Health Foundation (HERF), Richland, Washington, with support provided by Pacific Northwest Laboratory (PNL).** The USUR tissue program is the first systematic effort to contact and gain assistance of workers occupationally exposed to uranium and to seek and study uranium deposition in human tissue. This paper describes the program objectives, populations selected for study, procedures for enrolling workers in the program, types of tissue to be studied, and analytical procedures to be used. PROGRAM OBJECTIVES The objectives of the USUR tissue program are to: 1) Determine the distribution and concentration of uranium, if any, in the tissues of occupationally exposed workers. 2) Compare bioassay measurements of exposed individuals with the results of analyses of tissues obtained at autopsy. 3) Seek evidence of histopathologic changes related to any uranium deposition that may be present. 4) Conduct analyses of whole bodies, when available, to obtain more precise data on the uranium burdens, if any, in the body and organs, and especially the distribution in parts of the body, such as the skeleton, that are not usually accessible for sampling. 5) Develop data that will assist in evaluating a) the accuracy of current in-vivo measurement techniques, b) the propriety of existing regulations, and c) the adequacy of current protection programs. SELECTION OF POPULATIONS FOR STUDY Selection of the populations for the USUR tissue program was based on studies of the United States Uranium Registry described elsewhere (0c81). Potential populations for study were quite varied because of occupational exposure to different chemical forms of uranium, different levels and eras of exposure, varying ability to identify populations exposed in the past, and the general interest and cooperativeness of the populations. Nevertheless, certain identifiable populations emerged that were both willing to participate in the study and likely to provide useful information. Among the targeted populations are individuals who were exposed to uranium prior to the existence of modern exposure standards. These individuals are presently being enrolled by the USUR. The reliability of concentration and distribution data and the demonstration of effects are more likely in such individuals than in those who have received less exposure. However, the targeted groups also include individuals who Have been exposed to permissible levels of uranium, and reference employees in uranium plants who have not worked with uranium. The lesser exposures are of immediate interest because of the rarity of heavy exposures under current standards. The sampling of potentially exposed individuals with negative in-vivo measurement results may serve to check the adequacy of in-vivo measurements in the detection of deposition. [ENROLLMENT PROCEDURE] The enrollment of participants follows three patterns. 1) Presently employed workers are enrolled by the cooperative efforts of USUR staff with the medical, industrial relations, and/or health physics personnel of the uranium facility. Volunteers may be enrolled by medical personnel at the time of their physical examination or by health physics personnel at the time of a bioassay or lung counting procedure. Follow-up after initial contacts with currently employed uranium workers is carried out by on-site personnel, but subsequent follow-up is the responsibility of the Registry. 2) Retired workers are contacted by mail, with a covering letter from their former employer supporting the principles of the study but leaving participation in the program on an entirely voluntary basis. 3) An alternate method of enrollment is to obtain a USUR autopsy consent after death from the next of kin. As part of enrollment in the program, permission is requested for access to the individual's medical and exposure data, and a short occupational history is filled out. The autopsy permission agreement signed by the individual is for an initial period of 5 years; the USUR ordinarily seeks renewal of the agreement upon its expiration. While this agreement provides

Jan 1, 1981

By Bob Snashall

Introduction 1300 BC, Egypt. Pharaoh, the god-king, owned all things. He was the only mine operator. As the provider of all things, Pharaoh had great expectations of his officials who gathered the wealth. Pharaoh's official, the mine foreman, was at a gold mine site to see that royal expectations were met. For the official, it could mean a promotion to the good life here and to the godly life hereafter. When he checked the haul for sufficient progress, a lot was at stake. The miner wore a loincloth, perhaps a headband and, if he was a prisoner, ankle manacles. Only an oil lamp helped illuminate the hot, dusty blackness. A fire at the base of the quartz ore face competed for scarce air. The ore so heated crumbled at the prompting of copper wedges. Confined to a crouch, the miner tossed chunks of ore onto a rope-mesh which, when loaded, was drawn up and lugged out. On the surface, the gold was ground to dust. Then it was transported by donkey caravan to the royal depot. There it was weighed, recorded, and distributed to workshops. Many minerals mined Egypt had gold mines to the south in Nubia and to the east in the desert and Sinai. Indeed, gold underwrote Egypt's prosperity. With a constant gold supply, fewer hungry hands robbed burial crypts and tombs. Gold was sacred, "the flesh of the gods." The shiny metal financed the army that policed the desert mining routes and guarded the gold caravans from Bedouin marauders. Gold theft was an offense to the gods. Anyone caught with gold `in his lunchpail,' so to speak, could say goodbye to life, both in this world and the next. In addition to gold, Egypt possessed other mined riches that allowed the Egyptian civilization to flourish. From Sinai and Nubia came copper. So abundant was the red metal that it enabled Egypt to become the supreme power, before the advent of iron. Also mined were amethyst, turquoise, feldspar, jasper, carnelian, and garnet. These were used for the rich inlay work that distinguished Egyptian jewelry and cloisonne. But Egypt's most endurable and awesome material was its stonework - for statues and obelisks and in temples, tombs, and pyramids. Stone quarrying was a vast enterprise. One expedition boasted nearly 10,000 men. These included 5000 laborer soldiers, 130 skilled quarrymen and stonecutters, and - egads! - even 20 scribes. In addition, there were thousands of officials, priests, and officers grooms. There were even fishermen, to provide the multitudes with the catch of the day. Mining methods detailed In 1300 BC, quarrying techniques had changed little since the age of the pyramids some 1300 years before. At that time, in 2600 BC, limestone was locally quarried and fashioned into the blocks of the pyramids. A basic limestone mining method was tunnel quarrying. A ramp was built up to the face of a cliff. A monkey stage was then erected on a ramp. While standing on the stage, quarrymen carved out a rectangular niche in the cliff. The niche was large enough for a quarryman to crawl into. With a wooden mallet, he hammered long copper chisels along the edges of the niche floor to free up the back and sides of the block. The quarryman climbed out of the niche and removed the stage. He then carved out a series of holes in the cliff face for what would be the bottom of the block. The quarryman pounded wooden wedges into the holes. He watered the wedges until they were soaked. The water-logged wedges expanded, splitting the stone along the line of holes. The freed-up block was then levered down from the cliff. On the ground, the blocks were placed on sledges. Men pulled these to nearby water transport. Without block and tackle pulleys, paved roads, and wheels, this was no mean feat. Each block weighed an average of 2.3 t (2.5 st). Whenever possible, the quarrying was done directly from the surface. This "open cast" quarrying also involved using chisels

Jan 2, 1987

By T. P. Lim, J. A. Cherry, A. J. Vivyurka, R. D. Blair

The first phase of an investigation of the chemical composition of subsurface water and the hydrochemical processes in and near the abandoned Nordic tailings at Elliot Lake, Ontario, was conducted in 1979. The Nordic area, which covers 85 hectares with an average thickness of 10 m and overlies deposits of permeable glaciofluvial sand, contains 15 million tonnes of tailings with 4-7% pyrite. Nests of standpipe piezometers installed at three locations on the tailings indicate that seepage through the tailings is predominantly downward. The water table in the tailings slopes from the north side of the tailings dam, where it is within a metre of surface to the south side near the tailings dam, where it is approximately 10 m below surface. When they were deposited during the period from 1957 to 1968, the tailings contained pore water at a pH of about 8 resulting from neutralization treatment in the mill. At the three piezometer nests, the pH increases from 3 near the water table to 7 near the bottom of the tailings. The pore water in the tailings is gradually becoming acidic because of oxidation of pyrite in the zone above the water table. The concentrations of Fe, Mn, Ni, Co, Zn and Pb are high in the shallow acidic zone and much lower in the deeper neutral zone. Concentrations of 226Ra range from 30 to 230 pCi/L. At the three sites concentrations of major ions in the permeable sand a short interval beneath the tailings indicate 226Ra presence of tailings seepage, but Ra concentrations are only 2 to 8 pCi/L. Greater downward penetration of 226Ra probably occurs in the area closer to the tailings dam. Monitoring of a network of multilevel bundle piezometers installed in permeable sand that extends southward from beneath the tailings indicates the occurrence of a plume of tailingsderived water containing high concentrations of Ca, Mg, Fe and S04 that extend 400 m downgradient from the tailings dam. The total dissolved solids decline from 20,000 at the dam to less than 500 mg/L at the identifiable downgradient periphery of the plume. A much smaller but distinct plume of water with high heavy metal concentrations and 226Ra in the range of 10 to 130 pCi/L extends from beneath the dam for a distance of 15 m. The pH of the groundwater in this heavy metal-radium plume is 4 to 5, whereas elsewhere in zone of groundwater with high dissolved solids, the pH is above 6. The results of this investigat, suggest that the heavy metals an 226 Ra are relatively immobile in the perme-

Jan 1, 1980

By S. M. Moore

Each year more than 400 coal miners are injured (fatally and nonfatally) by rock falling from between or around roof supports. Many of these injuries can be prevented by the installation of roof screen. However, many coal mines are reluctant to use roof screen because of the added cost. The goal of this study was to determine the potential savings in workers' compensation (W C) premiums that can be achieved due to a reduction in rock fall injuries after roof screening. The WC rate-setting methods utilized by Illinois and Kentucky were investigated in this study. Using data obtained from the Mine Safety and Health Administration, national and state WC bodies and individual insurance companies (e.g., average cost per injury, loss cost rate, number of injuries per year, number of injuries preventable each year with roof screening), baseline mines (representing two mine sizes: 67 and 150 employees) were constructed with realistic ranges for estimates of injuries and WC premium costs. Using each state's actual WC rate-setting formulas, a sensitivity analysis was conducted to determine the total savings in WC costs after a three-year period. Annual savings in WC premiums ranged from $41,000 to $326,000 for the 67-person mine, and $96,000 to $843,000 for the 150-person mine. An economic analysis of the cost of a roof screening program at a 67-person mine was also conducted. The annual cost of a roof screening program here was estimated at $240,000. At this mine, the reduction in WC premiums alone could pay for the entire screening program.

Jan 1, 2010

By Wolfgang R. Wawersik, Charles M. Stone

INTRODUCTION Hydraulic fracturing was applied in horizontal drill- holes in the Salado salt formation near Carlsbad, New Mexico. Tests were performed approximately 650 m be- low surface to validate the design of a Waste Isolation Pilot Plant (WIPP) for the disposal of radioactive waste from defense activities of the United States. Hydraulic fracturing was performed primarily to determine whether the virgin in situ stress state at the WIPP site is isotropic and whether the magnitude of the virgin in situ stresses corresponds to the weight of the overburden. Beyond these limited objectives, however, measurements are now being reviewed to evaluate the usefulness of hydraulic fracturing in salt formations in general. Such measurements are desirable to determine stresses induced by mining and to monitor time-dependent stress changes around underground excavations in salt masses. Hydraulic fracturing was chosen because it is the only technique that can be readily applied in virgin ground far from existing mine openings. The method also appeared to yield directly the magnitude of the least compressive principal stress, regardless of the mechanical properties of the rock mass. A knowledge of the least principal compression was deemed sufficient to establish the isotropic character of the virgin in situ stress field under two conditions: (1) the Salado formation was mechanically isotropic, i.e., the fracture toughness did not vary normal and parallel to bedding so as to permit preferred fracture paths, and (2) pressure measurements were combined with determinations of the orientations of hydraulic fractures and, more generally, with observations of the characteristic hydraulic fracture patterns. The latter was possible by locating the drillholes for hydraulic fracturing along the axes of drifts that were mined several months later. The success of general stress measurements by hy- draulic fracturing of salt subjected to anisotropic stress conditions was and remains less certain. While the smallest in situ compressive stresses should still be equal to instantaneous shut-in pressures, there may not exist any unique relationships between the greatest and intermediate in situ principal stresses on one hand and the primary breakdown and fracture reopening pressures on the other. This problem would be a consequence of stress relaxation due to salt creep around any drillhole in which hydraulic fracturing tests are conducted. Moreover, lab- oratory measurements suggest that creep and stress relaxation could be so rapid that it might not be possible to interpret the measurements for anisotropic stress conditions even if hydraulic fracturing were performed as soon after drilling as possible. The primary objectives of the stress measurements in the WIPP were met by combining hydraulic fracturing tests with finite element analyses of the effects of salt creep, observations of hydraulically induced fracture pat- terns, and laboratory hydraulic fracturing tests on oriented core. In situ hydraulic fracturing tests in long drillholes were preceded by several trial tests to establish the best test methods and to demonstrate that salt could be fractured. The resolution of anisotropic stresses and special problems, e.g., in the identification of instantaneous shut-in pressures, are currently being addressed in hydraulic fracturing tests in a mine pillar, in tests following different delays between drilling and fracturing, and in tests with different frac fluids. This paper will review the main results of the analyses above and of one group of hydraulic fracturing tests at different distances from existing excavations. The paper will conclude with a brief description of the geometry of the hydraulically induced fracture patterns. Because of space limitations, it will be assumed that the reader is familiar with hydraulic fracturing or will first consult other sources (e.g., Cleary, 1980; Gronseth and Detournay, 1979; Zoback and Haimson, 1983; Nemat-Nasser et al., 1983; Kim et al., 1984)

Jan 1, 1986

By Donna L. Boreck

Introduction Coal has been mined in the western United States since the mid-1800s. The resource is abundant in the West. Despite recent economic conditions, it remains an important part of the nation's energy warehouse. Many of the seams that comprise the western resource base are thick, some exceeding 61 m (200 ft). Multiple, closely-spaced seams or thick coals that split into two or more minable seams are also common. Historically, subsurface development recovered from less than 15% to 50% of these deposits (Matson and White, 1975). This resulted in a substantial loss of inplace reserve due to the limited height that could be mined using the available technology. In an effort to increase future recovery from such deposits, the US Bureau of Mines analyzed the potential for incorporating three thick-seam mining methods into western coal mines. These methods are high-face longwall, multislice longwall, and longwall caving. They were developed in Europe and are being used to mine thick coal seams in Europe and Asia. One aspect that will determine their success in the US is the geology of the individual deposit; namely, are the methods adaptable to the geologic conditions characteristic of western coal? This paper analyzes some of the geologic factors that would affect development of thick or closely-spaced seams using these mining methods. Occurrence of thick and closely-spaced seams Since mining conditions and methods differ around the world, a thick seam is often defined as a seam whose total thickness cannot be extracted using the available equipment. In the western US, the maximum extractable mining height is about 4.3 m (14 ft). For this research, a thick seam was defined as a minable seam 4.6 m (15 ft) thick or more. Since one of the seams is usually lost using standard mining methods, closely-spaced or split seams were also addressed. These were defined as two minable seams 1.5 m (5 ft) thick or more separated by 9 m (30 ft) of interburden. Thick or closely-spaced coal seams are common in the West. Their importance has not been documented due, in part, to the methods used to measure resources. In the past, all reserves greater than 1.1 m (42 in.) thick for bituminous seams and greater than 3 m (10 ft) thick for subbituminous seams were grouped together and were differentiated only by rank. Further differentiation by seam thickness or minability was not done. For this reason, reserve figures for thick or split seams are difficult to determine. By analyzing available published resource data, Ferm and Muthig (1982) addressed thick seam resources. They estimated a resource of 90 Gt (99 billion st) of coal (16% of the Western resource base) in seams 4.6 (15 ft) or greater at less than 610 m (2000 ft) overburden. Yet, a later study of Pierce and others (1982) on one thick seam deposit in the Powder River Basin estimated a resource of 103 Gt (113 billion st). Pierce (1986) reported that the overburden reached a maximum of 731 m (2400 ft). The resources in this single deposit exceeded the 90 Gt (99 billion st) total projected for the entire western US by 12.7 Gt (14 billion st). Due to such differences, this author chose to look at reported occurrences of thick or closely-spaced minable seams in place of resource figures, to verify the need to research thick seam mining in the West. An occurrence was defined as an isolated site where a thick seam or two or more closely-spaced seams were reported. The occurrences were derived from published material - mostly mine descriptions, measured sections, coal exploration logs, and oil and gas logs. Due to time constraints, the reserves associated with each occurrence were not determined. All data points were random. The analysis was limited to occurrences in Colorado, Wyoming, and Utah. Although source material available to the author was limited to general coal summa-

Jan 3, 1988

By Michael Murphy, Gamal Rashed, Brent Slaker

Many underground stone mines in the United States operate on multiple mining levels. The effects of interburden thickness, pillar offset between mining levels at various depths of cover, and in situ stress conditions are poorly understood. Pillar offset is the degree to which pillars are not columnized or stacked. In this study, calibrated numerical models are used to examine limestone pillar stability in these conditions. Also, the critical interburden thickness required to minimize the interaction between mining levels is explored. The model results show that there is interaction between numerous factors that control the stability of stone pillars in multiple-level conditions. The highest degree of local pillar instability occurs when pillar overlap ranges between 10 and 70 percent. The highest degree of stability occurs when pillars are stacked.

Jan 1, 2023

By Michael MacMillan

AI is already changing how mining companies are seen, judged, and trusted. This paper discusses “reputation engineering,” a structured way to build trust in a world where AI systems summarize and evaluate everything a company says publicly. Using mining examples, emerging AI search trends, and real communication work, the paper explains why the next 12–18 months are a crucial window for action. It also provides a practical framework to help mining companies improve their visibility, accuracy, and consistency across both human and machine channels.

Feb 22, 2026

By Duport. P. J.

INTRODUCTION In February 1981, the Canadian Institute for Radiation Safety (CAIRS) initiated a routine Personal Alpha Dosimetry service for personnel of the Canadian uranium mining industry. This service is based on the use of the [Personal Alpha Dosimeter] developed by the French Atomic Energy Commission (CEA). The origins of personal alpha dosimetry and its rational are briefly described. Technical and organizational aspects of a routine personal alpha dosimetry service are outlined in this paper. HISTORICAL BACKGROUND International recommendations (1) and Canadian regulations have established Maximum Permissible Exposures (MPE) for each source of radiation exposure. Uranium workers in mines and mills are exposed to external radiation ( [y] rays) and to internal radiations ( [B] and [a] particles) which are delivered to the respiratory track by airborne alpha emitters (Rn and Th daughters and Long Lived Dust). To date, dosimetry for uranium workers has been performed by area monitoring/collective dosimetry. In North America the concentration of radon daughters is routinely measured by grab samples taken at the work place and by on-site gross alpha counting. The concentration of potential alpha energy is then calculated (usually by Kusnetz method) and expressed in Working Levels (WL). The time spent by each worker at a given work place is determined from his time sheets and used to calculate the individual monthly exposures to airborne alpha emitters, which is then expressed in Working Level Months (WLM). The uncertainties attached to such a procedure are obvious even in the case of frequent grab samplings and can be expected to lead to an underestimation of individual doses. Among fifteen possible sources identified in a mine situation, (2) four may stretch the standard deviation of the measurements' distribution, nine may lead to an underestimation and two may lead to either an underestimation or to an overestimation. To improve this situation, in 1971 the Atomic Energy Commission began studying the use of personal alpha dosimeters to determine individual exposures from the airborne alpha emitters encountered in the uranium industry environments. Criteria for a Personal Alpha Dosimeter In order to minimize the difficulties encountered in determining exposures received by uranium workers, the CEA in co-operation with the Atomic Energy Control Board of Canada (AECB), has developed a set of criteria for personal alpha dosimeters. Exposures may be determined easily and accurately using this criteria. Autonomy The dosimeter must operate for at least 10 to 12 hours. Excess time spent in the mine or in the facility may possibly be related to an accidental situation causing unusual levels of radioactivity. Since the dosimeter may be needed in non-underground settings where a cap lamp is not used, full autonomy is desirable. Maintenance, Periodicity of Reading In order to complement other dosimetry systems, the personal alpha dosimeter should be read monthly when the filter should also be changed. Routine air flow checks can be made according to local conditions (e.g. diesel loading). Radioisotopes Identification Since the exposure unit (WLM) is based on the concentration of potential alpha energy in the air, the personal alpha dosimeter should be capable of identifying each short lived alpha emitter included in the calculation of the WI, and WLM. Permanent Exposure Record Three points may be considered here: 1. In many countries, lung cancer in uranium workers is a compensable occupational disease. In some instances, compensation is awarded when it can be proven that the worker has received an exposure above a certain limit. The present uncertainty of the individual exposure makes the compensation procedure difficult. 2. By design, a personal alpha dosimeter must representatively sample all airborne particles, ranging in size from the unattached fraction to the upper limit of respirable aerosols (0.001 to 5 µm). The dosimeter must offer minimal resistance to the penetration of these aerosols. While the mining/ milling environment presents harsh conditions which may accidentally contaminate the dosimeter, it is important to be able to distinguish these cases of contamination and still obtain accurate readings. 3. A dependable dose register is most valuable for further epidimiological studies. The dependability of such a data base increases with the possibility of a second assessment of the dosimeters' reading (filter, film).

Jan 1, 1981

By Earl M. Frizzell, Ronald W. Utley

Introduction For the past several years the Bureau of Mines has undertaken research to develop alternative mine material haulage systems that will improve mine productivity, reduce haulage costs, increase mineral availability, and conserve energy. One basic concept being pursued is a means for crushing primary blasted rock to a size that can be reliably and economically moved on conveyors. During 1980, mine requirements for this equipment and the capability of existing technology to satisfy these requirements were assessed. The study continued into 1981 on the basis that satisfactory equipment for this purpose needed to be developed. The objective of the 1981 program was to perform an engineering design of such a system. A mine survey and equipment development study was conducted in cooperation with the open-pit mining industry by GATX through its subsidiaries, Fuller Co. (a crusher manufacturer) and GARD Inc. (GATX's research and development center), under contract to the USBM. Results of a mine survey in the copper, iron, and molybdenum industries indicated that a universal crusher system could be developed even though details might vary from mine to mine. The crushing characteristics, including tonnages desired, were similar in all surveyed mines. Most mines had similar views about the desirability and frequency of crusher movement. The perceived need for in-nit crushing varied between copper and iron mining. The copper mines felt a greater sense of urgency for introduction of such a system, probably because typically greater vertical haulage is found in copper mines than in the iron mines. All mine personnel surveyed were interested in portable crushers and almost universally preferred gyratory-type primary crushers. They preferred large crushers, citing inherent reliability and durability, and the machine's relative freedom from clogging. Large size minimizes the number of crushers and conveyors needed. Analysis of desired crusher system characteristics (Table 1) fixed the crusher size at 1.5 x 2.3 m (60 x 89 in.). The 1.5-m (60-in.) feed-opening dimension could accept the largest material normally fed to the crusher, and the 2.3-m (894n.) head diameter could fulfill the 3.63 kt/h (4,000 stph) capacity requirement. Selected System Description The comments and analyses discussed in the contract report led to the design and refinement of a movable crusher system. This system represents the consensus of more than 100 mine workers and almost 50 construction and engineering company personnel. It is supported by more than 50 years of experience in crusher manufacturing. Although many alternatives are available for any given system, and selections are likely to be dictated by varying site conditions, the selected system is believed to be appropriate for most projected requirements. Final design would be dictated by mine operators during evaluation for approved projects. Detail design of the feed and discharge systems were not included in the study because they are mine specific. The basic system (Fig. 1) consists of a 1.5 x 2.3-m (60 x 89-in.) gyratory crusher on a structural steel frame. The frame is used as an operating platform for the crusher and as a carriage to move the crusher. The control tower includes a crusher lubrication room, a hydraulic equipment room, a switch gear room, a room for the motor control center, and a control room for operating the crusher and maintaining the equipment. The auxiliary system consists of a hydraulic rock hammer and pedestal crane with 74-t (80-st) and 14-t (15-st) hooks. Figure 1 also shows stairs, railings, and a work platform. Feed and discharge arrangements are shown on Fig. 2. The system shown is for mildclimate applications. Additional enclosure and heating would be needed for extremely cold weather conditions. Crusher A 1.5 x 2.3-m (60 x 89-in.) gyra¬tory crusher was selected for this application. This size can handle the typical throughput of 3-3.63 kt/h (3,500-4,000 stph) desired by the mines and the 1.3-m (50-in.) material to be crushed without difficulty. The crusher is the heaviest

Jan 4, 1983

By Robert E. Blackett

To help local governments plan for long-term supplies of sand and gravel for buildings and infrastructure, the Utah Geological Survey is assessing natural aggregate resources (mainly sand and gravel) in the St. George, Utah area. Whereas much of the aggregate in the study area is adequate for road-base construction, high-quality aggregate for Port- land cement concrete is limited due to geology, and also due to land administration and environmental issues. Virgin River alluvial gravels southwest of St. George are generally poor sources of aggregate because they contain clasts of soft sediments and soluble minerals derived from Mesozoic rocks. Elsewhere, widespread secondary calcium carbonate (caliche), as rinds and matrix in gravel, lower the quality of material. Young river-terrace deposits in Fort Pearce Wash, and along the Virgin River above St. George are the primary sources of sand and gravel for the area. Alluvial fans on the east side of the Beaver Dam Mountains, and on the west side of the Hurricane Cliffs are also sources of aggregate. Stream-channel deposits associated with "inverted valleys" north of the towns of St. George and Hurricane, Utah, and older pediment gravels on the west flank of the Beaver Dam Mountains may be future sources of high-quality aggregate. Accelerated urban development during the 1980s and early 1990s has caused St. George and other communities in Washington County to encroach upon sand and gravel pits originally located in rural, undeveloped areas. Urban residents often view nearby sand and gravel operations as undesirable neighbors and pressure elected officials to close the pits. Pit closures effectively eliminate the remaining reserve from the sand and gravel resource base of the area. More- over, resources may also be lost as new homes are built upon undeveloped sand and gravel deposits. The St. George area is particularly susceptible to loss of aggregate resources because high-quality natural aggregate for construction is already in short supply. Moreover, most of the land in the study area is administered by the U.S. Bureau of Land Management and subject to various land-use restrictions. Much of this public land cannot be developed for mineral materials, such as sand and gravel, because of federal regulations protecting rare plants and animals, archaeological sites, and important watershed areas. Of particular concern to resource development are areas designated as Areas of Critical Environmental Concern, Wilderness Areas, and Wilderness Study Areas. Protection for Endangered Species is a critical element in all planning efforts. Those species present in the St. George area or in areas near- by include: the desert tortoise, southwestern willow flycatcher, Virgin River chub, and dwarf bearclaw poppy. As the St. George area is depleted and/or deprived of sand and gravel resources through extraction and urbanization respectively, construction costs will rise unless conservation measures are instituted. To avoid the loss of aggregate resources, the UGS is pursuing a program of data gathering and field surveying of surficial deposits to determine tonnage and quality of the remaining sand and gravel resources. The resource information from this program will help land planners, developers, and government officials make informed decisions concerning urban development throughout the region.

Jan 1, 2001

By R. W. Carn, D. Vincent

During the next 15 years, US coal production is expected to double, with the increased production evenly divided between the East and the West. Along with greater production, coal export markets should increase dramatically from East, West, and Gulf Coast ports. The annual overseas export capacity of US coal-loading terminals is expected to rise from 147.1 Mt (162.1 million st) in 1981 to a minimum of 278.1 Mt (306.6 million st) in 1985, according to the US Maritime Administration. Increased coal production and use will lead to more development of import and export terminals, a vital link in the coal transportation chain. With continually escalating capital costs and the competitive markets that the terminals will serve, a well designed and efficient terminal is necessary. This article begins a two-part series that presents concepts presently used in coal export terminal design. Part I looks at site selection factors and equipment needs, while Part II will examine environmental considerations in building a terminal as well as typical capital and operating costs. The world is nearing the end of the oil era. In a few years oil will not be available to sustain the growth rate and increasing standard of living we have known in our lifetime. The big question is what energy era are we moving into? With the decline of readily available oil reserves and rapidly increasing prices, many countries are trying to switch to alternate energy forms. While intensive efforts to find new oil reserves continue, alternate energy sources such as natural gas, coal, synthetic fuels, nuclear, hydroelectric, solar, and wind power are being developed. Recent indications are that coal is expected to bridge the energy gap over the next 25-30 years until the technology and economics of the alternate energy forms reach satisfactory levels. Use of coal for energy is receiving strong attention due to its long-term availability (200-300 years minimum), relative ease of development, and its low cost per unit of power produced. By the year 2000, it is expected that 25% of world energy supply will be met by direct coal combustion and possibly another 5-10% by synthetic fuel from coal. Coal's expanding share in the world energy market, along with an increase in coking coal requirements, will result in a large increase in the world's seaborne coal trade. Recent statistics and projections for the future are shown in Table 1. This phenomenal development rate includes increases in both coking and thermal coal requirements. Because of the rapid increase of seaborne coal trade during the last 10 years and the even greater projected increase of trade to 2000, various sectors of the coal industry are faced with enormous technical challenges and huge investments in equipment, land, transportation systems, and port facilities. Very large bulk terminals are under development throughout the world. Latest surveys indicate that there are about 30 new coal export and import terminals under consideration and at least 30 existing terminals have expansion programs planned or underway. With the high cost of borrowed capital and rapid inflation rates there is great emphasis on new planning and design techniques to minimize capital and operating costs of coal transportation systems. Terminals A total coal supply system can be considered to consist of one or more mines; a train, barge, truck, or other haulage system; an export terminal; a fleet of bulk carriers; a receiving terminal; and possibly, local inland distribution networks that include barges and railways. Terminals, though only a small link in the total transportation system, play a key role in overall system efficiency. At ports or inland distribution centers, terminals act as transportation links bringing trains, ships, barges, or trucks together for cargo transfer and temporary storage. A well-designed terminal can provide maximum independence between two modes of transportation and optimum freedom for intermodal interference. A terminal acts as a buffer between the two transportation modes by providing sufficient storage capacity so a ship need not wait for its cargo on, for example, a train-by-train basis, but can load immediately from the ready stock. Similarly, a train need not wait for a ship to unload its contents but can dump immediately into storage. A terminal also can be used to properly mix various types of coal to satisfy a buyer's requirements. Consider the relative value of various production and transport segments for a typical steam coal

Jan 6, 1983

By Sunfangze Liu, Xumeng Chen, Yongjun Peng

Efficient separation of naturally hydrophobic gangue, such as talc and carbonaceous materials, from sulphide minerals in flotation is becoming a significant challenge in many processing plants due to the need to process lower grade and more complex ores. Considering the hydrophobic nature of these gangue minerals, prefloating them at the beginning of the flotation circuit in the absence of collectors and rejecting the concentrate is a common practice to remove these gangue minerals. However, sulphide minerals, such as chalcopyrite, can become naturally floatable due to surface oxidation and therefore may float together with naturally hydrophobic gangue. In this study, a new approach was developed to maximize the prefloat of hydrophobic gangue while minimizing the loss of sulphide minerals. The new approach introduced a prefloat cleaner stage where sulphide minerals recovered to the prefloat concentrate was depressed and separated from other hydrophobic gangue minerals at a low pulp potential with the prefloat cleaner tailing fed back to the main sulphide flotation circuit. Flotation tests were conducted using chalcopyrite and organic carbon. The results show that the proposed approach can significantly reduce the loss of valuable minerals in the prefloat flotation without affecting the reject of naturally hydrophobic gangue. Applying pulp potential control at the prefloat cleaner flotation stage also reduces the consumption of reducing agents while minimizing the effect of reducing agents on downstream rougher flotation.

Nov 23, 2020

By QUANWEN PU, Yonggang Wei, Bo Li

The spent catalyst is one of solid waste resources, which contains precious metals with considerable recovery value. However, due to the different compositions of catalyst, the recovery process is difficult. Most of the catalysts are supported by alumina which has good alkali-soluble properties, while platinum is insoluble in alkali. Platinum is enriched by pressurized alkaline solution which is a kind of hydrometallurgical process. The novelty of this study is the separation of alumina and platinum from the spent catalyst and the enrichment of platinum in the solid phase. In this study, a leaching process using NaOH as leaching agent under high temperature and pressure was employed. The effects of alkali concentration, leaching temperature, leaching time, and liquid–solid ratio on leaching efficiency were investigated. The optimum process conditions were determined as follows: alkali concentration of 240 g/L, temperature of 200℃, pressure of 1.20 MPa, time of 1 h, liquid–solid ratio of 5:1. It was concluded that the leaching rate of alumina can reach 99.69%. The content of platinum increased from 0.1749 to 21.12%, the loss of platinum was 3.39%, and the enrichment ratio of platinum increased by 120.82 times. The leaching kinetics of alumina was studied based on the shrinkage core model. The activation energy of alumina leaching is 19.53 kJ/mol, and the leaching process is controlled by chemical reaction. Pressurized alkali solution can not only efficiently leachate alumina from the spent catalyst and enrich platinum, but also reduce the reaction time and increase reaction rate.

Feb 11, 2022

By L. Zhou, J. P. Rider, T. W. Beck, J. A. Organiscak, Y. Zheng

The National Institute for Occupational Safety and Health’s Office of Mine Safety and Health Research recently developed a series of models using computational fluid dynamics (CFD) to study the gas distribution around a continuous mining machine with various fan-powered flooded bed scrubber discharge configurations. CFD models using Species Transport Model without reactions in FLUENT were constructed to evaluate the redirection of scrubber discharge toward the mining face rather than behind the return curtain. The following scenarios are considered in this study: 100 percent of the discharge redirected back toward the face on the off-curtain side of the continuous miner; 100 percent of the discharge redirected back toward the face, but divided equally to both sides of the machine; and 15 percent of the discharge redirected toward the face on the off-curtain side of the machine, with 85 percent directed into the return. These models were compared against a model with a conventional scrubber discharge, where air is directed away from the face into the return. The CFD models were calibrated and validated based on experimental data and accurately predicted sulfur hexafluoride (SF6) gas levels at four gas monitoring locations. One additional prediction model was simulated to consider a different scrubber discharge angle for the 100 percent redirected, equally divided case. These models identified relatively high gassy areas around the continuous miner, which may not warrant their use in coal mines with medium to high methane liberation rates. This paper describes the methodology used to develop the CFD models, and the validation of the models based on experimental data.

Jan 1, 2015

By Zhiyuan Ma, ZUI WANG, Xincheng Hu, Jianwei Cheng, Junhong Si, Yi Qin, Gao Ke

The harmfulness and destructiveness of mine gas explosion is quite serious, which is the first of all kinds of coal mine disasters. Once the gas explosion occurs, it could instantly generate explosion shock waves, high temperature airflows, and disaster gases, which cause the destruction of underground ventilation structures, disruption of airflow direction in a mine ventilation network, and miners’ deaths by asphyxiation. As an important barrier to prevent the spread of accidents caused by the gas explosion, the safety and reliability of the mine seal affects the mine’s production and the health of workers. Due to the demands of coal mine production and transportation, a large number of longitudinal and horizontal roadway networks have been developed underground. There are not only many flat and straight roadways in the complex roadway network structure, but also a large number of different types of bifurcations and bends. Once a gas explosion occurs, the impact of the blast waves on the mine seal after passing through the complex path is difficult to determine. Therefore, this paper establishes a numerical model of bifurcated and cornered roadways and studies how the impact of gas explosion shock waves on the structure of sandwich mine seal after passing through different types of bifurcated and cornered roadways. The damage state of sandwich mine seal is analyzed. The researching finding could prevent and manage the spread of gas explosion disaster, and improve the disaster mitigation of mine explosion accident.

Jul 11, 2023

By W. J. Schlitt

Introduction A major uncertainty in assessing the potential for heap-and dump-leach projects is how to determine the extraction-rate curve for the recovery of the mineral values from coarse ore. Such material could either be run-of-mine (ROM) or primary crushed ore. The problem with field testing coarse ore, especially for new projects, is the large scale and extended leach times needed to accurately determine the final extraction-rate curve. At least 5 x 103 to 5 x 104 t of representative ROM ore are typically required for a copper test heap, and much more is often used. Kennecott, for example, recently constructed a 0.9 Mt (1 million st) ROM test heap at the Bingham Canyon Mine in Utah. In such coarse ore operations, the ultimate level of extraction will require a leach cycle that can extend from several months to a few years. Quite often, project development schedules do not provide the luxury of mining such large quantities of material or running such long tests. Instead, test data are usually limited to results from column leach studies on relatively fine ore, often with a top size that does not exceed 25 mm. Maximum leach times are also short, typically less than a year before an initial decision is needed on project viability. Proposed method One approach to estimating the recovery from a coarse ore leach is to assume that the leach solution will have some ultimate penetration distance into the rock. Then, the final level of mineral extraction in this "leached rim" will equal the ultimate level of extraction identified in various testing programs. Obviously, if the radius of a given rock fragment is less than the penetration distance, that fragment will be fully leached at the end of the operation. With larger rock sizes, the percent recovery will fall off as the size increases and the fraction of unpenetrated rock mass increases. Such an approach sounds simple but is likely to be complex when applied to a real project. For example, the penetration distance will be a function of both the rock characteristics and the effective length of the leach cycle. The important rock characteristics include rock porosity, the degree of internal fracturing and the mode of mineral occurrence. With regard to the latter, penetration is likely to be greater if the leachable mineralization occurs on fracture surfaces or in veinlets, as opposed to fine grains uniformly disseminated throughout the rock mass. An estimate of penetration distance may be derived from column or heap tests by noting the depth of solution penetration into the larger rock fragments after three, six and 12 months of leaching. While the penetration rate is ore specific, something on the order of 10 to 20 mm/y may be appropriate for competent, primary copper (chalcopyrite) ore. For gold in tight quartz, the rate may be about the same. Copper oxide ores and gold that is hosted in a more porous rock matrix are likely to have penetration rates that are at least two to three times higher, and an even higher rate should be appropriate for uranium hosted in sandstone. As noted above, the length of the effective leach cycle is likely to be measured in years. On this basis, the ultimate penetration distance (dp) would vary from less than 50 to several 100 mm when a particular ore is leached to exhaustion. Several sets of mathematical manipulations are necessary to convert a rock size distribution and corresponding value of dp into an estimated extraction-rate curve. The first step is to break the ROM size distribution down into intervals and then calculate the radius for the mean rock size in each interval. This is shown in Table 1 for rock sizes up to 1.75 m (about 6 ft) in diameter. The next step is to calculate the volume of unleached core and the fraction of rock that is leached. This is done for the following three values of dp: 25, 100 and 250 mm. Results are shown in Table 2. The third step is to select the ultimate level of recovery that will be achieved in the fraction of material that is effectively leached, i.e., the outer zone that is penetrated by the leach solution. This is clearly a site-specific factor that can only come from metallurgical test results on representative ore

Jan 1, 1998

By Louis J. Trostel

Refractories, called the Hidden Industry by The Refractories Institute, provide the temperature and chemically resistant linings for the multitude of vessels that are used today in high temperature processes ranging from giant iron-making blast furnaces to industrial incinerators to petrochemical cracking units to the smallest furnaces used by the jewelry trade. While these refractory linings are absolutely essential to the proper functioning of these processes, the linings are almost always on the inside of the vessels and thus hidden from the eyes of the public. Refractories are defined in Standard C71 of the American Society for Testing and Materials (ASTM), titled "Standard Definitions of Terms Relating to Refractories," as "nonmetallic materials having those chemical and physical properties that make them applicable for structures, or as components of systems, that are exposed to environments above 1 000°F (538°C)." This criterion of ability to withstand exposure to environments above 538OC is the critical distinction separating refractories from other ceramics, fibers, and coatings applicable only at lower temperatures. In addition to resisting temperature, the refractories must with- stand abrasion, chemical and slag attack, resist thermal shock, and carry sustained structural stresses at high operating temperatures for as long as years. The most common methods of classifying refractories are by forming method and composition. First are the traditional formed and fired bricks, batts, and the associated multitude of special shapes. These refractories are usually formed at room temperature, dried, and then fired to high temperatures to develop the final properties desired. Other formed refractories are chemically bonded or resin or tar bonded. To complete the formed refractory matrix, some fired refractory bricks and shapes are subsequently impregnated with resins or tars. Physical dimensions of these formed refractories are classified into two general categories: standard sizes and special shapes. The standard sizes are those in common usage in the United States described in ASTM Standard C909 "Dimensions of a Modular Series of Refractory Brick and Shapes" based on the 38 mm basic module as described in ASTM Standard C861 "Determining Metric Dimensions of Standard Series Refractory Brick and Shapes." The most common size is the "9-inch straight" 228 x 114 x 64 mm (9 x 4 ½ x 2 ½ in.) of the "2 ½ -inch" series or 228 x 114 x 76 mm (9 x 4 ½ x 3 in.) of the "3-inch" series. Bricks of nonstandard sizes are referred to as special shapes. Their use is required in specific furnaces for particular requirements or applications. In addition there is the large and growing group of refractories sold as unshaped mixtures that may be formed into their final shapes at the application site. These unshaped mixtures are sometimes referred to as specialties. The mixtures are prepared to be hydraulically cast, plastic or wet rammed, dry rammed, or dry vibrated into their final shapes. Some are designed to be projected through nozzles (or gunned) onto their forms. The refractory mortars used to lay up the shaped refractories fall into this group referred to as specialties. Prepared refractory grains also are sold. These are usually calcined or fused materials that are subsequently crushed and sized to a variety of specific carefully controlled grain sizes or size distributions. These grains include calcined fire clay, dead burned magnesite, periclase (fused magnesia), dead burned dolomite, kyanite, tabular and fused aluminas, fused zirconia, fused mullite, and calcined alumina-magnesia spinel. The primary composition groups of refractories are the 1) alumino-silicate refractories made largely from fire clay and high- alumina materials, 2) silica refractories made from quartzites and ganisters, and 3) basic refractories from magnesia (or magnesite) and chromia (also referred to as chrome ore), alone and in combination, and dolomite. Special refractories include carbon, zirconia, zircon, high purity alumina, mullite, silicon carbide, silicon nitride, and some borides. In addition there are groups of special insulating refractories.

Jan 1, 1994