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By A. Bleier, E. D. Goddard, R. D. Kulkarni

The influence of the collector structure on the amine (RNH2, R=alkyl) flotation of quartz has been systematically investigated using a modified Hallimond microflotation cell. This study focuses on the effect of varying chain length, branching in the alkyl group R, polyfunctionality of the collector molecule, and alkyl substitution on the nitrogen atom. In addition, adsorption at the solution/air interface and electrokinetic results for the solid/solution interface were employed to further elucidate the role and the interrelation of the structural parameters in the flotation process. Estimates of an electrical double layer attractive energy between the bubble and the solid particle are made.

Jan 1, 1976

MULSIM/NL and La Model are displacement-discontinuity programs which have been used quite successfully in the past to provide detailed structural models of various single and multiple-seam mining situations. However, both of these programs required fairly complex details of the mine plan and topography to be manually input into the program, often taking several days of work. Recently, automatic mine and topographic grid generation capabilities have been added to the MULSIM/NL and La Model programs, reducing the input time requirements from days to hours. This paper discusses the algorithms used for the automatic grid generation and demonstrates the utility of the new capabilities with a case study.

Jan 1, 2003

By George W. Luxbacher

The inclusion of a new term "planned subsidence" in the Surface Mining and Reclamation Act of 1977 has lead the U.S. coal mining industry into a new aspect of engineering: subsidence planning. Subsidence planning has necessitated the integration of geotechnical, civil, and mining engineering disciplines. Although much of the foundation and application work has been completed, subsidence planning is still a developing area. This paper discusses the application of subsidence planning in several brief case studies (a high voltage transmission tower, a sewage treatment plant, and a slurry impoundment), emphasizing the risk assessment aspects necessary to account for some of the uncertainties that may exist.

Jan 1, 1986

By Maria C. Jaime

Optimization of the rock cutting performance plays a major role in the tunneling operation. For a given set of tool and cutting environment, rock fragments may be chipped away in numerous manners and it is crucial to understand and be able to predict the dynamics at the cutter-rock interface. This paper presents rational and realistic 3D finite element modeling of rock cutting where typical computational issues ?such as contact formulation, appropriate rock material response, mesh distortion, and crack propagation? are overcome. Cutting forces and failure mechanisms in this finite element analysis are successfully validated by laboratory single-cutter scratching tests in sandstone.

Jan 1, 2010

By James A. Barclay

Results of investigations over approximately the last 10 years by the Bureau of Mines and others on technology for recovering cell-grade alumina from domestic nonbauxitic resources such as clay, anorthosite, alunite, dawsonitic oil shale, coal ash, and coal shale are reviewed. It is concluded that, although many innovative techniques have evolved from this effort and have been shown to be technically feasible, energy and economic considerations will prevent any of the processes from competing with the Bayer process under normal conditions. Based on the Bureau's experience, some variation of clay-HCI leaching, HC1 gas sparging crystallization has the greatest potential as an alternative to the Bayer-bauxite process.

Jan 1, 1984

By Paul M. Godlewski, Roberto J. Guardia

Hundreds of highway and rail transportation tunnels in North America have been rehabilitated in the last two decades. Tunnel rehabilitation is usually dictated by an owner’s need to extend the working life of the structure, but typically has the added benefit of increasing capacity (greater speed or increased clearance) and enhancing safety. Not only has rehabilitation proven to be an economical alternative to new tunneling, but it can be safely accomplished by working during short work “windows” without taking the tunnel out of service. This paper presents methodology for planning and designing of tunnel rehabilitation projects, and case histories illustrate successful excavation and support technologies.

Jan 1, 2003

By S. E. Phillipson

Slickensides cause potentially hazardous ground conditions in underground coal mines. Investigations by MSHA’s Roof Control Division indicate that many slickensides represent bedding plane faults and drag folds that formed during coal bed decollement, a style of faulting that occurs in shale units above coal seams. These slickensides formed in response to regional tectonic compressional stresses, rather than by differential compaction, and are parallel to folds and thrust faults. This paper relates slickensides in the Appalachian, Illinois, and Black Warrior basins to the structural grain of the Appalachian Mountains, provides descriptions of slickensides, and advocates geologic mapping as a predictive method for identifying areas with adverse ground conditions.

Jan 1, 2003

By Daniel W. H. Su

The ability to predict the effects of longwall mining on topography and surface structures is important for any coal company in making permit applications and anticipating potential mining problems. The sophisticated finite element model described and evaluated in this paper is based upon five years of underground and surface observations and evolutionary development of modeling techniques and attributes. The model provides a very powerful tool to address subsidence and other ground control questions. The model can be used to calculate post-mining stress and strain conditions at any horizon between the mine and the ground surface. This holds the promise of assisting in the prediction of mining-related hydrological effects.

Jan 1, 1986

By John E. Anderson, Charles A. III Cravotta, Eleanora I. Robbins, E. Terrance Slonecker, Gordon L. Nord

Remote sensing is being tested for its ability to identify the severity of impact on streams that are subjected to contaminated mine drainage (CMD) in the Southern Anthracite Coalfield of Pennsylvania. Airborne digital multispectral video images of acidic and near-neutral CMD are being compared with field spectra, water chemistry, mineralogy, and microbiology. Spectral differences in the visible region of the electromagnetic spectrum are caused by the presence of schwertmannite in the acidic precipitates and ferrihydrite in the near- neutral precipitates. The acidic CMD averages 1,500 mg/L sulfate and has bacteria coated with schwertmannite; the near-neutral CMD averages 87 mg/L sulfate and has bacteria coated with ferrihydrite.

Jan 1, 2000

By G. Gregory Bryan

G. GREGORY BRYAN 33.1-DEFINITIONS 1. Engineering-Engineering may be defined simply as the science and art of making materials and energy serve the purposes of mankind. There are many fields of engineering, each with specialized branches. The basic fields are: civil, mechanical, electrical and chemical. Others may be specialized phases of one of these or a combination of two or more of them. Mining engineering, for example, requires a working knowledge of construction, mechanics, electricity, metallurgy, geology, accounting and economics, in addition to mining techniques or methods. Mining geology and geological engineering involve rocks and their structure and the forces that act upon them, and chemistry, physics and mechanics also are important to practitioners in these fields. Metallurgical engineering requires chemistry and physics but also a working knowledge of mechanics and electricity. The fields of engineering with which this section are concerned are those primarily involved in the finding, mining, processing and utilization of the natural mineral resources of the earth. They include geology, mining engineering and metallurgy, each with several specialized subdivisions. 2. The Engineer-The engineer is an individual who has had specialized training to qualify him to practice the profession of engineering. The term engineer is used here in its broadest sense to include such professions as certain applied branches of geology, physics and chemistry, whose practitioners may not have acquired engineering degrees but who, nevertheless, by training and experience in some area of the minerals industry, qualify under the above definition. To become a successful engineer, a person should have certain pronounced aptitudes and preferences, even before embarking upon the requisite course of training. Among the chief qualifications that the engineer-to-he must possess are: (a) a natural inclination toward mathematics and the physical sciences, and (b) intellectual and moral honesty. Among the qualities the student engineer must develop are: (a) independence of thought, open-mindedness, sound judgment and the ability to observe, deduce, plan and apply solutions, (b) the ability to articulate and convince, and (c) the ability to organize and direct the activity of others. These qualities are normally developed through sound education, followed by pro¬gressive experience. 3. The Consulting Engineer-The consulting engineer is an engineer primarily in the business of furnishing specialized advice and engineering services on an intermit¬tent, part-time or occasional basis to clients who do not have continuous or full-time need for those services. He must be able to make and take full responsibility for decisions and recom¬mendations where original thought is required. He should not be confused with the engineer, surveyor or technician who also may furnish services on an intermittent basis, such as regular periodic surveys, the overseeing of drilling programs and other such routine work as may require little original thought and planning. 33.2-THE PROFESSION OF CONSULTING ENGINEERING 33.2.1- QUALITIES REQUIRED A successful consulting engineer must first have, as a minimum, the qualities and preparation of an engineer. In addition, he must have sufficient knowledge,

Jan 1, 1973

By Dan Nilsson

INTRODUCTION There are many mine planning factors in sublevel caving, as in the other mining systems, which when varied can substantially alter mining costs and profit¬ability. In this chapter, the following four topics are addressed with regard to economic optimization of sub¬level caving: production planning, haulage level spacing, orepass spacing, and extraction cutoff. The costs used here are the right order of magnitude but since each mine is unique, they should be considered only as examples. The objective is to demonstrate the tech¬niques which can be used. In a real mine, the evalua¬tions must be done using actual costs and conditions. It is difficult to predict exactly the costs involved, and therefore it is valuable to perform a sensitivity analysis to evaluate the effect of making incorrect as¬sumptions. Since the mining industry is very capital intensive, the effect of the interest rate must also be studied. PRODUCTION PLAN FOR AN IRON ORE DEPOSIT Introduction The first and most important thing to do before de¬signing an underground mine is to establish a long-range production plan. Such a plan is necessary for all eco¬nomic evaluations and should provide information about the lifetime of the mine, how much ore and waste must be handled per year, how much development per year is required, etc. An example is given in the following section. Problem In an iron ore mine sublevel caving is used. The iron content is 42%, and the mine supplies a pelletizing plant with an annual capacity of 3 million t/a. The ore body is shown in Fig. 1. The length of the ore body is 1000 m and the width is 100 m. Each slice is 10 m high, and there are 10 m be¬tween crosscuts, each of which has an area of 20 m2. The density of the ore is 3.5 t/m3. The spacing between rings is 2 m, and the extraction is 100%. The iron con¬tent is 66% in the pellets and 6% in the tailings. The problem is to develop a detailed production plan. A typical sublevel caving sequence is shown in Fig. 2. Solution Amount of ore per meter of crosscut: 20 X 3.5 = 70 t. Number of crosscuts per slice= 100. Length of crosscuts per slice: 100 X 100 = 10 000. Ore from crosscuts per slice = 10 000 m X 20 m2 X 3.5 = 700 000 t. Number of sublevel caving rounds per slice: 10 000/2 = 5000. Area for each blast in sublevel caving: 10 X 10 - 20 = 80 m2. Amount of ore per blast: 80 m2 X 6 m X 3.5 = 560 t. Extraction: 100% (see extraction curve Fig. 12). Loaded ore per blast: 75% or 420 t. Loaded waste per blast: 25% or 140 t. Total: 560 t. Total amount of rock from each blast in the sublevel caving: Ore: 560 t of which 2 X 70 = 140 t from develop¬ment work and 420 t from sublevel caving. Waste: 140 t. Total: 700 t. The ore needed per year is 3 X (66-6)/42-6 5 mil¬lion t. Necessary number of blasts per year= 5,000,000/ 560 = 8929. Distance to develop per year = 8929 blasts per year X 2 m per blast 17 858 m. Total amount/year: Ore from development work 17 858 m x 70 t/m = 1.25 million t/a Ore from sublevel caving: 8929 blasts per year X 420 tons per blast = 3.75 million t/a 5.00 million t/a Waste rock dilution: 8929 blasts per year X 140 tons per blast 1.25 million t/a Total amount to hoist 6.25 million t/ a The amount of ore from development work will in¬crease a little if the horizontal drift is placed in the ore body and not in the footwall. In this example the ore loss in 20%, and the waste rock dilution is also 20%. After taking the ore loss into account, the lifetime for 100 m of the ore body will be:

Jan 1, 1982

By William D. Cassity, Batric Pesic

Introduction Gold mining on an industrial scale is becoming increasingly dependent on low-grade refractory sulfide ores that contain large amounts of pyrite (FeS2) and arsenopyrite (FeAsS). Gold is usually present as sub-microscopic particles dispersed evenly throughout the sulfide matrix. Without some form of preoxidation, efficient recovery of this finely disseminated gold through cyanide leaching is hampered because the surface of the sulfide particle becomes passivated and impervious to penetration of the cyanide. Traditional methods for recovery of refractory gold from arsenopyrite ores include roasting and autoclave leaching. A modem alternative method is to preoxidize the mineral using the bacterium Thiobacillus ferrooxidans prior to attempting to re- cover the gold through cyanidation. The bioleaching of arsenopyrite ores presents a problem in the mobilization of large quantities of arsenic, present both in the +5 and +3 ionic states. Dissolved arsenic species pose a dual problem: they are -1 to the bioleaching process itself through inhibition of bacterial activity, and they also pose an environmental threat. The arsenate anion (AsO2) has been shown to be more toxic to living organisms, including Thiobacillus ferrooxidans, than the arsenate anion (AsO43) (Collinet and Morin, 1989). There is some disagreement as to the fate of arsenite in bioleaching solution. Mandl, Matulova, and Docekalova (1992) reported that dissolved arsenite concentration was constant during a long-term bioleaching study of chalcopyrite using Thiobacillus ferrooxidans. Torma and Oolman (1992) report that Thiobacillus ferrooxidans has been found capable of oxidizing As3+ to As5+. The solid products of arsenopyrite/pyrite bioleaching include jarosite (MFe3(SO4)2(OH)6 where M can be K+, Na+, NH4+ or H3O+), scorodite (FeAsO4.2H2O), ferric hydroxides, and ferric hydroxysufites (Carlson et. al., 1992; Van Breemen, 1982; Lazaroff et. al., 1982). The particular species formed is a function mainly of pH. Little study has been performed on the effect of solid products on bioleaching. Pesic and Kim (1993) showed that Thiobacillus ferrooridans cells served as a nucleation sites for jarosite particles, which grew rapidly and eventually killed the cell. Most bioleaching studies are performed at low pH&apos;s (<2.0) under uncontrolled conditions. By studying bioleaching of arsenopyrite under controlled pH conditions and at higher pH&apos;s (2.0 to 3.0), the distribution of iron and arsenic between the solid and liquid phases can be controlled, and their effects studied. The objective is to develop a bioleaching process that operates at high efficiency while keeping amounts of dissolved heavy metals low. Pesic, Stohok, and Torma (1993) demonstrated the usefulness of this approach in the leaching of cobaltite (CoAsS) concentrates. Materials and Methods The focus of this study was the bioleaching of arsenopyrite ore by Thiobacillus ferrooxidans. The parameters studied included the effects of pH, and the effects of arsenate (AsO43 and arsenite (AsO2). pH studies included pH set initially but not controlled, and pH controlled continuously with NaOH or LiOH. Bioleaching results from two strains of Thiobacillus ferrooxidans with different adaptation histories were compared. Bioleaching experiments were conducted in 150 ml stirred glass reactors using a strain of Thiobacillus ferrooxidans adapted to arsenopyrite ore, supplied by the U.S. Bureau of Mines. These bioreactors were innoculated with stock cells grown on arsenopyrite ore for 14-16 days. Typical elemental breakdown for this ore was: As, 9.90%; Fe, 18.60%, Stot, 9.26%; Suifide, 7.50%; SiO2, 25.90%. Mineralogically, the ore contained both arsenopyrite and pyrite. Bioleaching efficiency was measured by withdrawing bioleach solution samples at selected intervals, filtering, and measuring the amount of dissolved cobalt, iron, and arsenic by atomic absorption spectroscopy. Because the precipitation of iron and arsenic at higher pH&apos;s would lead to unreliable results, a 100 mg tracer of cobaltite concentrate from the Blackbird mine in southern Idaho was added to solution. Dissolution of the cobaltite tracer was used to indicate bioleaching of the bulk arsenopyrite ore. Cobalt has been shown to be stable in solution even when iron and arsenic were precipitated (Pesic, Storhok, and Tonna, 1993). The pa-

Jan 1, 1995

By J. Le Gac, P. Zettwoog, J. Brenot, J. W. Piechowski, J. C. Nenot

INTRODUCTION The components of the radiological hazard in uranium mines are: 1. exposure to radon and its short-lived daughters 2. exposure to long-lived ore dust 3. external irradiation With usual ventilation rates, the dosimetry relative to the first component only requires to know each miner&apos;s exposure to a potential energy of short-lived radon daughters, (EXP[a]). Presently in France, (EXP [a]) is determined monthly from radon concentration measurements. This is an indirect method based on ambient monitoring. After technical improvements achieved during the last years it is now possible to measure (EXP [a] directly with an individual portable monitor developed by the CEA. The purpose of this study is to analyze and compare monitoring results obtained by these methods. DETERMINATION OF INDIVIDUAL MONTHLY EXPOSURES TO[ a] POTENTIAL ENERGY OF SHORT-LIVED RADON DAUGHTERS [(EXP a) amb, by ambient monitoring] In this case, it is necessary to know individual monthly exposures to radon, (EXP Rn). As mentioned by Pradel, J.,et al., 1976, (EXP Rn) is estimated for each worker from informations both on radon concentrations (CRn) measured as described by Pradel, J., 1962 and times (t) elapsed in the various workplaces. (EXP Rn) is obtained by summing up the elementary radon exposures CRn x t relative to all the workplaces where the miner was successively engaged during the month. (EXP Rn) values are then corrected by an average equilibrium factor of 0.17 deduced from many measurements and taken as representative for all the French mines. Finally: [ ] (EXP [a])ind are obtained directly with an individual portable instrument described by Chapuis, A.M., et al., 1972. Samples of the air breathed by the miners are pumped continuously and radioactive aerosols containing the short-lived radon daughters are trapped on a filter. The principle of the measurement lies on track-etch detection of the a particles emitted by the deposited material. The technical aspects have been described by Duport, P., et al., 1980a, Mouden, A., et al., 1980 and Chapuis, A.M., 1980. As mentioned by Duport, P., et al., 1980, fiability is good. Indications on measurement reproducibility are given by Pradel, J., et al., 1977. PRESENTATION OF THE STUDY The underground miner population of a French mine was followed during a period of 8 months (October 1978 - May 1979). Monthly exposure values for each miner determined either by ambient or individual monitoring, (EXP [a])amb or (EXP[ a])ind, had been previously collected by Le Gac, J., 1980, who had also defined miner categories according to the characteristics of the work performed. Four groups are considered here, the "whole mine" and three categories as described in table 1. Data on the individual monthly exposures,(EXP [a])ind and (EXP [a])amb, were taken again to analyze the individual cumulative exposures instead of the "month by month" values. The influence of the mean habits of each individual on its surveillance results is thus taken into account. Each miner has a cumulative exposure as an individual of the "whole mine" and as an individual of each category in which he has worked. The results obtained by ambient or individual monitoring were analyzed (characteristics and adjustements of the distributions) and then compared (description of the discrepancies observed). Obviously, such a study is possible only if all the exposures are cumulated on a same duration. As miners have not worked during the same number of months in the course of the data collecting period, one must considerer the individual average monthly exposures, [(EXP a)]amb and[ (EXP a)]ind, instead of the individual cumulative exposures, which is typically the same thing. To be sure that the exposure values of a given group (A, B, C or D) are representative, only those miners really present for at least 3 months in this group were taken into account. Finally, each individual of the "whole mine" and of the various work categories has a pair of average monthly exposure values : [(EXP a)amb] determined by ambient monitoring and[ (EXP a)ind] deduced from the track-etch recording of the individual portable monitor. Let a miner be engaged during n months[ (3 < n < 8)] in a group, the following relations then apply : [ ]

Jan 1, 1981

By Richard L. Bullock

At the beginning of the previous chapter, variation in the types of room-and-pillar stopes were briefly identified so that the reader could begin to understand the extensive application of this mining system. These same variations will be reviewed and elaborated upon in this chapter while discussing the methods by which the material is extracted. FULL-FACE SLICING BY DRILLING AND BLASTING When the entire mineralized thickness is taken in one pass of mining, it is known as full-face slicing. There is no mineral of economic value intentionally left either in the floor or the back. Theoretically, there is no limit as to how high the face could be in a single pass. But there are practical limitations to equipment size, pillar stability, and control of local loose slabs on the face, roof, and pillars that dictate current practices. The range of face height of the 15 typical room- and-pillar mines covered in Table 8 (Sec. 1, Subsec. 7) from the Dravo report, 1974 was 1.7 m (5.5 ft) to 9.8 m (32.1 ft), the average being 5.2 m (17.1 ft). In the Gaspe copper mine, Murduchville, Que., the drill jumbos were constructed with an extendable tower for full-face mining up to 15.2 m (49.9 ft) if necessary. However, a more common practice at that mine is to take the ore in multiple slices (Hall, 1959). In many metal mines the normal practice when starting stoping is to drive a single development drift into the ore zone a distance that will allow at least four or five rooms to be opened on each side of the initial drift. Since this opening will probably serve as the main haulage drift, it should be kept as straight and as level as possible. As to the other criteria for the stope developments, it might be well at this point for the mine planner to review the remarks presented earlier (Sec. 1, Subsec. 7) on drifts, entries, and crosscuts for production. After the initial drift is driven, it should be slabbed to the full room width if it is not that wide already. Next, the future pillars should be marked on the ribs and the rooms driven between them. If the ore extends beyond the length of the initial drift and is to be mined in the normal sequence of mining, care should be taken in a random room-and-pillar system to see that the extension of the initial roadway remains straight. If pillar "spotting" is left to the shift foreman or to the miners, they probably will not realize the importance of maintaining a uniform pillar line next to the roadway, and of keeping the line straight. Invariably, a pillar will end up in the middle of the future drift extension resulting in a "dogleg" in a possible main haul- age road. While this error is not catastrophic, too much weaving in and out of pillars will certainly slow down future production haulage and needlessly increase costs. Face or Breast Drilling and Blasting Practices To drill and blast the initial advances into the rock, usually some form of cut pattern must be used. In room-and-pillar mining, when a pattern is drilled in a face and that face is the only exposed surface to which all of the rock must be broken, the pattern is known as a "face round" or "swing." The face to which the rock must be broken is referred to as a "free face." Thus the most common way to advance a room into virgin rock with only one free face is by drilling swings. If, after breaking around a pillar, taking down back, or taking up bottom, there is a second (or third) face exposed, a group of holes drilled nearly parallel to a free face and breaking to it is known as a "slab round" or "slabbing." In some mining districts it is also known as "slashing." Obviously, because two faces exposed offer less resistance to fragmentation than a single face, slabbing requires less drilling and fewer explosives than required to break the same tonnage with the same degree of fragmentation with swings. Therefore, it behooves the driller (or supervisor) to plan the combination of

Jan 1, 1982

By Kathleen Kreiss, A. James Ruttenber

INTRODUCTION Early in the morning of July 16, 1979, there was a breach in the earthen retaining dam of a tailings pond at the United Nuclear Corporation&apos;s (UNC&apos;s) Church Rock uranium mill. The acidified liquid and tailings slurry spilled through the damaged portion of the retaining wall into an arroyo that is a tributary to the Rio Puerco river system. The Rio Puerco runs through Gallup, New Mexico, and eventually crosses the New Mexico-Arizona border (Fig. 1). On its way to Gallup, the Rio Puerco and its tributaries pass through land with a checkerboard pattern of ownership, with portions owned or leased by the Navajos, individuals, the Bureau of Land Management, and the State. In terms of tailings liquid volume (3.6 x 108L; 94 million gal), the UNC spill ranks as one of the largest. The mass of solids released in the slurry (10.0 x 105 kg; 1 100 tons) appears to be close to the median for accidents of this kind, however [U.S. Nuclear Regulatory Commission (NRC), 1979]. The UNC first opened its Church Rock uranium mill in 1977 on land adjacent to acreage belonging to the Navajo tribe. The mill, which is next to the UNC Church Rock mine, is located approximately 16 km (10 miles) northeast of Gallup, New Mexico (Fig. 1). Gallup, a town of 18 000 people, is the closest population center. The region surrounding the plant site is sparsely populated by Navajos, at a density of approximately 5.8 persons/km2 (15 persons/sq mile). The UNC mill and mines employ approximately 650 persons, and the adjacent Kerr-McGee uranium mine employs about 300. The UNC mill normally processes 3.2 x 106kg/day (3 500 tons/day) of uranium ore, depositing the acidified tailings slurry in a series of three earthen holding ponds. The tailings ponds are located east of the pipeline arroyo that feeds into the Rio Puerco approximately 2.4 km (1.5 miles) from the southernmost tailings dam. The liquid portion of the tailings slurry evaporates in the ponds; hence, under normal conditions, there is no surface flow from the holding ponds to the arroyo. Both runoff from the plant site after heavy rains and possible seepage from the tailings ponds may deliver radionuclides to the arroyo-river system, however. The dam across the southernmost tailings pond was considered to be in keeping with the state of the art. However, the New Mexico Environmental Improvement Division (NMEID) had warned UNC about dangers of locating the pond over a heterogeneous geological formation. The state Engineer&apos;s Office approved of the site only after UNC agreed to strict design criteria. Others have pointed to dangers of constructing earthen dams for impoundment of uranium mill tailings (Carter, 1978). Causes of the dam break were multiple: the UNC mill filled the tailings pond to a level that exceeded permit criteria; the tailings pond was lined improperly; the dam was constructed using clay that was compacted excessively, resulting in cracking and subsequent seepage; and the unstable substrate beneath the dam permitted differential settling. The UNC Church Rock mine has continuously released dewatering effluent into the pipeline arroyo at a rate of 88.3 L/sec (1 400 gal/min) since 1968. Before 1975 this effluent was not treated; after 1975 it received precipitation treatment for removal of Ra-226. Radionuclides are also released into the river system through the dewatering of the Kerr-McGee uranium mine 1.6 km (1.0 mile) north of the UNC mill. During usual mining operations, approximately 227 L/sec (3 600 gal/min) are released into the pipeline arroyo and subsequently into the Rio Puerco. The Kerr-McGee mine began continuous release of dewatering effluent in January 1972. In 1974 Kerr-McGee began Ra-226 precipitation treatment of its dewatering releases, but NMEID data indicate that treatment has often been incomplete. The effluent from both mines has been responsible for transforming the downstream portion of the Rio Puerco from a sporadically dry riverbed into a continuously flowing stream and has contributed to the current levels of background radiation along the river system (Table 1). This paper will summarize the postspill monitoring efforts and relate the assessment of this spill to the general question of evaluating the health effects of nuclear fuel-cycle wastes. The data pertaining to the measurement of radionuclides in the Church Rock environment and the radionuclide concentrations in animals will appear in forthcoming reports. CHURCH ROCK HEALTH EFFECTS ASSESSMENT APPROACH The initial health effects evaluation involved identifying the radionuclides that were released into the river system from the tailings pond. Table 1 lists the State of New Mexico maximum permissible radionuclide concentrations for liquids released to unrestricted areas, the typical tailings liquid concentrations, and postspill river water concentrations. The tailings liquid contained comparatively high levels of Th-230, Ra-226, Pb-210, and Po-210--all of which, according to postspill river water samples, had exceeded the state maximum permissible concentrations (MPC) at one time or another. After the radionuclides in the tailings were identified, pathways through which humans could be exposed were clarified. Environmental monitoring data were then used to quantify the important pathways of human exposure. Water samples were collected from the river, from test wells dug near

Jan 1, 1981

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

9.1 INTRODUCTION There are approximately 70,000 abandoned coal mines nation¬wide, which is about 35 times the number of underground coal mines presently operating. The US Bureau of Mines estimates that there are over 8 million acres of undermined land due to the extraction of coal, metals, and nonmetals. Subsidence has affected more than 2 million acres, and more than 99% of the subsidence is related to underground coal mining. There is reason to believe that some of the remaining 6 million acres of the undermined land have a high potential to subside. The expansion of housing, highways, commercial structures, and other facilities has required the use of many areas that are underlain by abandoned coal mines, and this growth will continue. Many subsidence problems are derived from the collapse of aban¬doned underground coal mines. There are numerous abandoned mine workings in the anthracite fields of northeastern Pennsylva¬nia, in the Appalachian bituminous fields, the Illinois Basin, the Rock Springs, Wyoming area, and other areas of the United States (Gray et al., 1976). Various room and pillar patterns of mining have been used in the dipping anthracite seams and the nearly flat-lying bituminous seams with considerable variation in the per¬centage of coal extracted. The progressive deterioration of pillars, mine floors, and mine roofs after long exposure to air and water may later result in the collapse of strata over the mine entries, the crushing of the remaining coal pillars, or the bearing failure of the mine floor beneath the coal pillars. Subsidence then results as the collapse reaches the ground surface in the form of differential strains, depressions, cracking of the ground, and sinkhole devel¬opment. Subsidence over active longwall mines, which occurs concurrently with mining or is completed within a short period following coal extraction, has been studied extensively over the past decade. On the other hand, subsidence over abandoned coal mines receives little attention by the researchers, mainly because it is difficult to predict and takes place decades after mining has ceased. The techniques of investigating the subsidence events over abandoned coal mines are similar to those employed for active mines except that at the outset, it is necessary to determine whether or not the subsidence events are mining-related (Chugh et al., 1986; Cummings and Singh, 1986; Peng and Hsiung, 1986). This calls for the identification and confirmation of abandoned mine workings under or near the affected surface structures. Gen¬erally old mine maps, if available, are acquired, the surface bore¬holes are drilled for confirmation of the accuracy of the mine maps and determination of the potential for continued subsidence. Sub¬sidence monuments are established and periodic surveys con¬ducted to determine the amount and trends of surface movement in and around the affected surface structures; surface boreholes are used to investigate the integrity of the underground structures (i.e., roof, coal pillars, and floor) by TV camera. They are also used for monitoring the vertical and horizontal movements of the subsur¬face strata by Sondex (FPBX) and inclinometer (PFBI), respec¬tively, for determining the continuity of the subsidence events. Tape extensometers, crackmeters, etc., are used to monitor the development trends of major cracks in the structures or on the ground. These data are used to identify the causes of the subsidence events. Finally, abatement methods are selected to stabilize the structures. It must be noted that most subsidence events over abandoned coal mines are reported after the fact, and investiga¬tions are begun some time after reporting, that the subsequent measured movements are generally much smaller than that at¬tained in the active mines, and that due to lack of knowledge about the damage conditions and the exact location of the abandoned mine workings with respect to the affected surface structures, the precise causes of surface subsidence or surface structural damages are in most cases very difficult to identify. 9.2 TYPES OF SURFACE SUBSIDENCE According to Gray et al. (1977), after examining 354 incidents of subsidence above abandoned mines in the Pittsburgh metropol¬itan area, the subsidence features have a mean diameter (i.e. the average of long and short dimensions) from less than 1 ft to 1600 ft, with 84% less than or equal to 15 ft; the subsidence features have a depth ranging from less than 1 ft to 48 ft, with 89% less than 25 ft; 66% of the subsidence features are deeper than they are broad. Nearly 59% of the subsidence features occur with overbur¬den less than 50 ft thick and 81% less than 100 ft thick. No subsidence features occur with overburden thicker than 450 ft (Fig. 9.1). Occurrence of subsidence incidents varies from imme¬diately to more than 100 years after mining (Fig. 9.2). In analyzing the characteristics of approximately 3000 chim¬ney subsidence features along the Colorado Front Range, Matheson and Eckert-Clift (1986) examined historical aerial photographs on 4- to 14-year intervals between 1937 and 1967 and found that the majority of observable surface subsidence features occurred within 30 to 40 years after mining. According to Gray et al. (1977), the most prevalent subsi¬dence features over abandoned mined land are sinkholes, with depths of more than 3 ft, and troughs or sags, usually less than 3 ft deep. Sinkholes are steep-sided pits, while troughs are shallow depressions much wider in area than sinkholes. A. SINKHOLE SUBSIDENCE A sinkhole is caused by the collapse of a mine roof that works its way upward. If it is not arrested during the process it will eventually reach the surface and emerge as a sinkhole. The process is governed by the thickness and character of the overburden, the width and height of the mine openings. Sinkholes are usually 3 to 20 ft deep and may be 2 to 40 ft in diameter, although most are fewer than 16 ft across (Gray et al., 1977; DuMontelle and Bauer, 1983). Newly formed sinkholes have steep sides with straight or bell-shaped walls. At times, they appear to be conical in profile with the apex upward. If the topsoil collapses, the top portion will widen to form an hour-glass shape. Sinkhole subsidence usually occurs over abandoned mines less than 165 ft deep (Hunt, 1979). Matheson and Eckert-Clift (1986) found that chimney sink¬holes are likely to occur when the ratio of overburden thickness to mining height (h/m) is less than 4 to 5. When h/m is between 5 and 10 to 11, the potential occurrence of chimney sinkholes decreases rapidly. When h/m is more than 10 to 11, less than 10% of the mine openings that collapse will induce sinkholes on the surface.

Jan 1, 1992

By M. M. Botz

Cyanidation tailings which are disposed in a surface impoundment experience a loss of cyanide due to natural degradation, frequently reducing the cyanide concentration to very low levels. Quantifying cyanide losses in terms of pond geometry, local weather conditions and feed solution chemistry has been largely empirical though in many cases mining operations rely on surface impoundments to reduce cyanide to below an internally regulated concentration or an effluent limitation. To permit a more quantitative evaluation of cyanide losses in an impoundment, a computer simulation was developed to estimate the losses of free, WAD and total cyanide due to dissociation, photolysis and volatilization. Results of the model are compared against data collected for a North American tailings impoundment during 1998.

Jan 1, 1999

By J. MCGUIRE, E. J. Haas, B. Demich

Near misses can enhance awareness of the potential causes of injury and prompt safety management initiatives. Although most companies require near-miss reporting, the value of these reports and how they influence subsequent actions or controls to reduce on-the-job risks are unclear. Researchers at the U.S. National Institute for Occupational Safety and Health (NIOSH) conducted a case study with an aggregates company in which near-miss reports were analyzed at each of their locations over an entire quarter during the summer of 2018. Within that quarter, workers recorded 249 near misses. The analysis provides insight into ways that risk-communication and management programs can be improved to reengage workers and their situational awareness on the job.

By Qingming Feng, Bo Huang, Dongbo An, Jinhong Zhang

A study of the application of mine tailings as precast construction materials through alkali activation was carried out, focusing on efficiently activating mine tailings, reducing alkali consumption, decreasing curing time and improving compressive strength. Test results showed that unconfined compressive strength of 40 MPa was achieved. In addition, to elucidate the geopolymerization mechanism of mine tailings, microscopic and spectroscopic techniques including scanning electron microscopy with energy-dispersive X-ray (SEM/EDX) analysis, Fourier-transform infrared (FTIR) spectroscopy and X-ray diffraction (XRD) spectroscopy were used to investigate the microstructures and elemental and phase compositions of the reaction products. The findings of the present work provide a practical method for applying mine tailings as precast construction materials through alkali activation.

By H. J. Chen

In recent years, it has been recognized that the high horizontal stress is an important factor that contributes to a cutter roof. When the immediate roof is thinly bedded, the presence of a high horizontal stress may first cause failure on the bedding planes or weak planes, and then a cutter roof develops as the individual stratum crushes and buckles progressively layer by layer. In this study, the stress distributions in the bedding planes of the immediate roof are analyzed for an entry-pillar system in the multiple longwall panels by using the finite element method. Then the roof failure is analyzed based on the rock failure criterion and the stress distributions.

Jan 1, 1998