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By Deepak Malhotra, William F. Riggs, Edward C. Dowling

INTRODUCTION Mining companies are In the business of making an acceptable profit on their Investments, It became obvious in the recent recession that economic profitability/viability depends on being the lowest cost producer In order to compete on a global basis. The burden of achieving this objective falls on the mineral manager. He must strive to produce a salable product in an environmentally acceptable process while continuously Improving the profit margln.(1,2) Metal prices fluctuate on a regular basis. They directly Impact revenue and hence operating philosophy. The mineral engineer needs to develop an optimum operating strategy to meet his overall objectives, One would expect that, In the modem era of complex computer models, it would be easy to develop an economic model to maximize profits and/or cash flow. Discussions with mill managers Indicate that one mine/corporation's optimum operating strategy is not necessarily optimum for another. This paper approaches the problems from basics "Economics 101" and Illustrates the complexity of developing a universal model which can be applicable to all mines. Case examples are used to illustrate the need for better communication to avoid the basic economic fallacy “that optimization of segments leads to optimization of the whole system.” TECHNICAL MODELS Technical models are a prerequisite to the development of economic models because they provide the necessary Input data to these models, The flow of material from a mine through the sale of the finished product is shown In Figure 1. The ore Is mined and milled, and the concentrate can be stored for sale or smelted and refined In-house or at a toll facility. The finished product(s) are held In Inventory until sold either through existing contracts or in the spot market. [ ] Fairly reliable technical models have been developed for mining, milling and smelter operations. Mine plans are developed to provide a good Indication of the relationship between production and cut-off grade. It is easier to change surface mine plans than plans for underground mines. However, there are limits to maximum production that can be achieved with existing equipment. It Is possible that mine benches may not be developed to handle additional equipment, even If It was purchased to enhance production. More often than not, the cut-off grade Is lowered to Increase throughput goals.

Jan 1, 1993

By Bernard L. Cohen

INTRODUCTION The most serious occupational radiation episode in U.S. history was the exposure of uranium miners to radon daughters in the 1945-1968 time period. Although the maximum permissible dose at that time was 12 WLM (working level months) per year, average exposures were several times that figure, and ten times higher exposures were not uncommon. Starting in the early 1960s, mounting epidemiological evidence appeared for an excess incidence of lung cancer among these miners; up to 1974, there was an excess of 134 cases (159 observed vs 25 expected) among the 4000 in the group under study (NAS-1980). When this excess became apparent, the Federal Government took jurisdiction, greatly tightened enforcement, and in 1969 lowered the maximum permissible exposure to 4 WLM per year. These standards were met largely by a great improvement in ventilation. As a result of these measures, average miner exposures were reduced 5-fold from 1965 to 1968 and by another factor of 3 by 1970. In 1978, the 4 WLM/year maximum permissible exposure was exceeded by only 37 of 7500 underground miners, for only 16 wars it exceeded by more than 25%, and for only 8 was it exceeded by 502 (AIF-1980). The average exposures to the various groups are listed in Table 1. We see that the average exposure for all miners was 1.03 WLM, and for those who worked essentially full time underground it was 1.45 WLM. For purpose of later discussion, it will be convenient to choose a single value for average exposures. In view of the fact that employment situations and job categories are bound to vary over a lifetime of work, we take this to be 1.3 WLM/year. The purpose of this paper is not to dwell on history, but rather to address the question of whether or not this present situation is satisfactory. This question was considered recently by a study group under the auspices of National Institue of Occupational Safety and Health (NIOSH-1980) and it answered the question in the negative, concluding that the maximum permissible exposure should be substantially reduced. However, their arguments were incomplete and were lacking in perspective. The principal thrust of this paper is to provide some of the material overlooked. RISK TO URANIUM MINERS FROM RADON EXPOSURE In order to quantify the risk to uranium miners under present working conditions, it is first necessary to estimate the risk of lung cancer per WLM of exposure. The most straightforward way of doing this is to use the data on the group of uranium miners under study. These are listed in Table 2 for 8 exposure ranges and for the total group (NAS-1980). If all data are given equal weight, the risk is seen to be 3.5 x 10-6 per year per WLM. Much can be said in favor of using this value for the risk as dose independent; it is within one standard deviation (SD) of the observations in 5 of the 8 dose categories (as expected from the definition of SD), and within 2 SD for all 8, and for the three cases differing from the mean by more than one SD, two are above and one is below. However, the uranium miner exposures in the present situation will be far below those covered in Table 2, so it is worth considering the possibility that the risk differs from this at low doses. From the last column of Table 2 we see that the experience from all exposures below 600 WLM indicates a risk higher than the mean for the entire group by more than two standard deviations (the value 600 WLM in this comparison was deliberately chosen to maximize this deviation; for exposures below 360 WLM and 840 WLM, the excess is only 1.0 and 1.6 standard deviations respectively). In view of this tendency for the risk to be higher at lower dose, we will take the risk to be 5.0 x 10-6 per year per WLM in what we will refer to as model A. This risk estimate is based entirely on exposures much [higher] than those that will be experienced by

Jan 1, 1981

By P. Brooks, W. LaRochelle, J. M. Tranquilla

EMR's recently commissioned Pilot Plant in Fredericton, N.B. Canada, the latest and largest in the company's sequence of prototype and demonstration systems, is being used as a process development facility for mining clients worldwide. Fitted with two microwave reactors and 75 kW generators, this plant is designed as a two-stage continuous processor with a throughput in the 5-10 tpd range depending upon the material feedstream. This facility has been used principally to calcine pyrite and arsenopyrite gold concentrates although it can readily accommo­date carbonaceous materials as well as a wide range of other ore types. Engineering details of the plant and the results of the first nine months of operation are presented. The plant has proven very successful as an advanced process development facility as well as a basic production unit module.

Jan 1, 1998

By Donald S. Bubar

The Separation Rapids property, located approximately 60 km (37 miles) north of Kenora, Ontario, Canada, is host to one of the largest rare metal pegmatite deposits in the world. This pegmatite was named the Big Whopper by its discoverer, Dr. Fred Breaks of the Ontario Geological Survey, because of the exceptional length and width of the surface exposure. The property is readily accessible from a main, all-weather road via a network of secondary logging roads. The main line of the Canadian National Railway passes through the village of Redditt, just 40 km (24 miles) south of the property. The Big Whopper is one of the complex-type (petalite sub-type) class of rare metal pegmatites that are geochemically the most highly evolved in the spectrum of granitic pegmatites. Such deposits are economically important as resources for the rare metals lithium, tantalum, cesium, and rubidium. While complex-type pegmatites are found in many areas of the world, most are too small to be profitably mined, however, with an inferred resource in excess of 15 million metric tons (mt) (17 million short tons [st]), the Big Whopper is only the fourth deposit of its type in the world with the size required to be of major economic importance. The other three deposits, which are currently in production, are the Tanco mine in Manitoba, the Bikita mine in Zimbabwe, and the Greenbushes mine in Western Australia. The Big Whopper is situated in the eastern continuation of the Archean Bird River greenstone belt which also hosts the Tanco pegmatite, approximately 60 km (37 miles) to the west. The principal commodities identified in the portions of the Big Whopper pegmatite explored to date are petalite (LiAlSi4010) and rubidium-rich K-feldspar. These are industrial minerals with important applications in the glass and ceramics industries. The Big Whopper also contains substantial quantities of lepidolite, a lithium, rubidium-mica which is the principal ore mineral for rubidium metal. Tantalum and cesium occur in anomalous levels with the petalite mineralization, and the possibility of finding zones of high enrichment in these valuable high-tech metals elsewhere in the deposit is excellent. Such enriched zones are typical of pegmatites as highly evolved as the Big Whopper. To date, the geological mapping and diamond drilling work completed by Avalon have delineated the Big Whopper pegmatite system over a strike length exceeding 1.5 km (0.9 miles), over widths ranging from 10 m to 80 m (30 feet to 260 feet) in thickness, and to a vertical depth of close to 300 m (1,000 feet), where it remains open. The pegmatite system consists of a vertically oriented massive petalite pegmatite dyke striking 280 degrees that is flanked by amphibolites containing a swarm of narrower albite and petalite dykes which have all undergone intense deformation in a high strain zone resulting in folding and intense shearing. The Big Whopper exhibits a mineralogical zonation pattern which is complicated by the superimposed deformation that has both folded and stretched the deposit. Drilling has defined a dilute geological resource totaling 7.1 million mt (7.8 million st) grading 1.283 percent Li20, 0.346 percent Rb20, and .007 percent Ta205 over a strike length of 600 m (2,000 feet), and to a vertical depth of 250 m (820 feet), where it remains open. The petalite and rubidium-rich-K-feldspars contained in the Big Whopper all appear to be of superior quality. The grades are consistent with a petalite content averaging 25-30 percent and a rubidium-rich-K feldspar content averaging 15-20 percent. The remainder of the rock consists mainly of albite, several types of mica, and quartz. Accessory minerals include columbitetantalite, cassiterite, apatite, garnet, and gahnite. Market studies show that the deposit is located close enough to existing transportation infrastructure to access rapidly growing, major markets in the glass and ceramics industries, both in the northeastern U.S. and Europe. There are only three competing lithium minerals producers in the world with the Bikita mine in Zimbabwe being the only major producer of petalite. The other lithium deposits produce spodumene, a less desirable lithium mineral. Lakefield Research Limited has successfully designed a process to produce an ultra-pure petalite concentrate containing up to 4.65 percent lithium oxide, and as little as 0.014 percent iron oxide, low levels of soda and potash, and negligible amounts of other trace elements. With these specifications, Avalon will have an excellent quality product for glass-ceramics applications such as Corningware®, CERAN® stove tops, and other thermal shock-resistant products. Lakefield is presently designing the balance of the flowsheet to produce separate concentrates of rubidium-rich K-feldspar, albite, mica, tantalum, tin, and high-purity quartz. Initial test work indicates that the feldspar products will be of exceptional quality. The K-feldspar concentrates contain 11-12 percent K20 and 1 percent Rb20, while the sodaspar contains over 10 percent Na20, and both have very low iron (0.01 percent Fe203). A preliminary economic model shows that with an initial production rate of 90,000 mt (100,000 st) per year, expanding to 170,000 mt (190,000 st) per year over 5 years, the project is capable of generating over CDN $10 million per year in pre-tax cash flow by year 5 on a capital cost of approximately CDN $30 million.

Jan 1, 2001

By W. T. Walter

High-G centrifuges from five preparation plants were sampled to determine and quantify the various operating factors affecting performance. The operating factors that were considered include feed rate and percent solids, size of feed material, screen-basket type, hardness of coal and age of the flights. The performance parameters that were measured include product surface moisture, recovery and degradation of the coal. Regression equations with good significance levels were developed.

Jan 1, 1982

The Waste Isolation Pilot Plant (WIPP) is a geologic repository for disposal of transuranic waste and is operated by the US Department of Energy. The repository is located at a depth of 655 min the Permian age salt beds of the Salado Formation. The exhaust air from the facility is independently monitored by the Environmental Evaluation Group to verify compliance with regulatory release limits for radioactive materials. A May 1995 video inspection of the exhaust shaft identified fluid migration into the uppermost 25 m of the shaft. Since then, water has been observed flowing in from cracks in the shaft lining between 16and 25 m below surface level at an estimated rate of between 5 and15 m3/d. This water inflow has not only negatively impacted the equipment in the shaft, associated duct-work and monitoring equipment located at the top of the shaft, but also challenged efforts to verify compliance at the WIPP. This paper provides an analysis of these effects.

Jan 1, 2004

By H. H. Murray

Tactical R & D is research and development for a specific application or applications. The mineral scientist has to define and understand the physical and chemical parameters required for a specific or potential application. Then the researcher must know the structure and composition of the mineral and its physical and chemical properties. Can these properties be changed by processing particle size and shape modification, and surface chemical treatment? Examples of tactical R & D in the industrial minerals area will be described.

Jan 1, 2003

By Heriban Soto

Primary amines are strong collectors of apatites and they can selectively float sedimentary phosphates from calcareous ores. However, secondary and tertiary amines and tri-methyl ammonium salts were ineffective. The use of kerosene reduces the collector requirements because hydrocarbon is coadsorbed along with the amine on the mineral surface. The heats of reaction of aminium ions with such anions as phosphate, carbonate and sulfate, were measured in a microcalorimeter and discussed in terms of the solubility of the precipitates formed and their relationship with the adsorption mechanism of primary amines on phosphate minerals. Experimental evidence supports a view that primary amines are adsorbed through chemical interaction between the aminium and phosphate ions at the mineral surface.

Jan 1, 1984

By Neville G. W. Cook, John Kemeny

Utilizing the principles of elastic-brittle fracture mechanics, stress-displacement relationships are derived for an elastic solid containing a slip plane consisting of various configurations of collinear arrays of cracks. These models take into account the heterogeneous nature of slip planes as observed in the lab and in the field, and are relevant to the behavior of rock containing faults, joints, bedding planes, and fractures. All of our models exhibit slip weakening behavior, and show regimes of both stable (aseismic) and unstable (seismic) deformation. For given configurations of cracks and applied boundary conditions, our results show explicit ly the energy released through unstable crack growth. For instance, it is found that two equal length cracks growing together release more seismic energy than, say, a large crack growing into a small crack. Also, by allowing cracks to coalesce, our models show multiple stick-slip events without crack healing or other geochemical means.

Jan 1, 1986

By Ned M. Smith

The Mississippian Salem Limestone, from which dimension stone is quarried in Indiana, is principally a calcarenitic rock formed of fossil bryozoans, echinoderms, and specimens of Endothyra. Numerical associations of the fossils, binding materials, oölitically coated fossils, and voids were determined by counting points on the surfaces of 278 thin sections of the limestone. Median sizes, coefficients of sorting, and skewness numbers were computed from measurements made of constituents in these thin sections. Many other samples and numerous exposures of the Salem and contiguous parts of adjacent formations also were studied. The organisms that furnished the skeletal material for formation of the Salem Limestone were chiefly gregarious and communal. Salem rocks are formed of nearly whole skeletons, of dismembered or slightly damaged skeletons, and of skeletons thoroughly ground up, because currents in the Salem sea varied in their ability to move and damage these materials. Locally the sediments are of uniform grain size because of sorting by currents and because related organisms grew to similar sizes. Rocks of the formation contain different proportions of the fauna both because the animals lived in these proportions and because their skeletons were mixed together by currents. The proportional relationships between members of the fauna were constant locally and for short periods of time. Study and mapping of the faunal associations, numbers of oölitically coated fossils, bedding structures, and size parameters enables geologists to suggest sedimentologic histories and sequences in areas of various sizes. Interpretations based upon such areal studies are necessary for evaluation of the dimension-stone potential within the Salem Limestone.

Jan 1, 1966

By James A. Finch

The Gaudin lecture is devoted to fundamentals. Given my involvement with column flotation over the last 20 years, this is my topic. I have tried to identify spinoff basic research and provide a personal list of favorites. Two are selected for discussion here. They are single bubble rise and bubble surface area flux. The rationale for such work, as well as some of the findings, are reviewed. Support for fundamental work seems as volatile as the economy (and the link is probably not coincidental). I hope some utility to the minerals industry is evident in the text. Column flotation in its current form originated in the early 1960s. The method was invented by Boutin, patented by Boutin and Tremblay and pursued doggedly by Wheeler. Their aim was to overcome the problem of mechanical entrainment of fine particles in the water carried by the bubbles, which they were convinced contributed to lowering flotation concentrate grades (Wheeler, 1988; Finch and Dobby, 1990). The solution was a combination of columnar geometry, the addition of wash water into a deep froth and the maintenance of an excess of water to the tailings, as compared to the feed (a so-called positive bias). Success at Les Mines Gaspe, Quebec, on cleaning Mo in the early 1980s signaled the start of industrial acceptance, making the column one of the flotation machines of choice today. Over the next 20 years, column applications mushroomed to cover virtually all mineral commodities, as well as nonmineral applications such as soil decontamination, deoiling of water and deinking of recycled paper. In the wake, research groups are continuing to probe the fundamentals, conferences and workshops have been transferring the technology and specialist companies have emerged to service the demand. Innovations were triggered as problems arose (for instance, in gas injection systems) and novel flotation cells, much in vogue in the 1990s, had the door opened. Mechanical cell manufacturers have not been idle - there is nothing like competition to provoke a fresh look at an established business. Flotation-cell technology has probably not seen so much activity since the patent rush in the early days of flotation, which occurred around the turn of the century. For an industry often considered reluctant to risk innovation, this is not a bad record. A good idea will find a place, but patience (not always an innovators strong suit) and careful preparation are required. Two books (Finch and Dobby, 1990; Rubenstein, 1996), three symposia (Sastry, 1988; Agar and Huls, 1991; Gomez and Finch, 1996), and a series of reviews (Dobby and Finch, 1991; Finch and Dobby, 1991; Finch, 1995; Finch et al., 1995) document the period. Like many good ideas, column flotation has been overexposed, and there have been disappointments. In a recent example, columns performing scavenging duty were replaced by mechanical cells (Chong et al., 1998). The further applications stray from the original - to

Jan 1, 1999

By R. D. Rein

The Henderson mine is a large continuous panel caving operation in central Colorado. The orebody is roughly 760 m by 910 m and will be extracted from two production levels. Each level is divided into three north - south panels 244 m to 317 m wide. The general arrangement of a production level and its associated undercut drifts is shown in Figure 1. The production drifts (3.6 m by 3.7 m) are spaced 24 m apart center to center and have undercut drifts (3.4 m by 3.8 m) 17 m directly above them. The production drifts are interconnected with drawpoint crosscuts (4 m by 4.3 m) driven on 12 m centers. Drawpoint crosscuts receive substantial brow steel and concrete support prior to undercutting. The brow steel is a 15 cm by 46 cm wide flange beam with a 2.5 cm wear plate to box in the 'cave' side of the beam. The concrete is a minimum of 0.3 m thick and the batch design results in 27 500 kPa strength. The finished drawpoint opening is 3.7 m by 3.2 m with a choke area of 3.7 m by 2.4 m. (Figure la) Undercutting and bell formation is accomplished by blasting rings of 7.6 cm holes drilled from the undercut drifts. The rings have a burden of 2.0 m and six rings develop one complete bell.

Jan 1, 1985

By Toshio Fujiwara, Sumikichi Ito, Masao Kitamura

The Tennohji-Benten Giant Trunk Sewer Project has been under construction since 1974 in the eastern part of Osaka, the second largest city in Japan. The entire project is divided into 7 blocks, some of which have already been completed. This 8.8-km sewer tunnel, being driven through a densely populated area, was planned to be constructed by the shield tunneling method. This paper described shield tunneling performances and ground behaviors during the construction of Blocks 2 and 3. An earth pressure shield (dia., 6,750 mm) was employed in Block 3 tunneling. The ground was generally soft and the cover over the tunnel varied from 13 m to only 4.5 m, so that a considerable amount of chemicals was injected to stabilize the ground. A hand-digging type shield (dia., 7,250 mm) was used in Block 2 tunneling under compressed air. For 85 m of the total length of 1.6 km, the soil at the face was extremely soft, and therefore, the ground freezing method was additionally employed.

Jan 1, 1981

By Emily Muto

LiDAR technology has made it possible to map large areas of underground mines accurately and rapidly in three dimensions. MSHA Technical Support began implementing LiDAR technology in conjunction with the Agency’s Pillar Collapse Initiative, using scans to determine width-to-height ratios (w/h) of pillars adjacent to mass collapse areas and to conduct risk assessments for global pillar stability in U.S. stone mines. This work quickly expanded to include change detection, such as measuring many pillars to show which are deteriorating over time. Change detection has also been used in ground falls and squeeze areas to see if they have continued to propagate. LiDAR technology supports accurate data acquisition of unstable mining conditions from safe distances. For example, once the LiDAR scan has been processed, it can measure roof fall cavity heights and volumes with the user at a stable ground location. Some LiDAR scanners can collect geologic data, such as joint orientations.

Jun 25, 2024

By C. H. Rawlins

The total amount of ferrous slag generated in the U.S. is about one-third the amount of iron ore mined, thus ferrous slag can be considered a significant mineral body for processing. Steelmaking slag contains up to 30 wt.% metal of which only the >25mm fraction is commonly recovered. A grinding study was conducted of BOF, EAF, and LMF slags using the Bond Work Index method to determine grindability to 106µm. Work Index values obtained were 13.8-24.9 kWh/ST, depending upon the slag type and composition. Steel particles were magnetically recovered from the crushed and ground samples for size-by-size analysis, yielding up to 18 wt.% of the slag weight recovered as magnetic. The recovered material, as characterized by inert gas melting, was found to be 61-96 wt.% metallic.

Jan 1, 2008

By W. B. Greer, J. J. K. Daemen

Introduction The performance of seals in boreholes, shafts and tunnels is an important concern in mine safety, in the control of effluents from operating and abandoned mines and in the underground containment of hazardous wastes (Einarson and Abel, 1990; Mining Waste Study, 1988). Despite the clear need for effective seals, few documented test results are available concerning the performance of in situ seals. Some evidence indicates that horizontal or near-horizontal holes are especially difficult to seal properly (Kelsall et al., 1982). This research involved the installation, extensive hydraulic testing, removal and post-test examination of a 10-cm-long x 10-cm-diam cement grout seal in a near-horizontal borehole in limestone. The work was performed in an underground mine on Marble Peak in the Santa Catalina Mountains, AZ. A test hole was diamond cored in the recrystallized limestone of the Escabrosa formation. The hole is 33.24 m long, 10 cm in diam, and dips at 10.4°. The hole extends between two drifts and thus is accessible from both ends. Seal installation Surveys to assist in selecting the interval to seal consisted of core logging, video logging and constant-head injection tests using a straddle-packer unit. The location selected, 3 m from the lower end of the hole, had low hydraulic conductivity (3 x 10-10 to 3 x 10-9 cm/sec) and was free of significant fractures. The grout used, an expansive Class A cement, was provided by Dowell-Schlumberger, Tulsa, OK. A grout mass about 1.5 m in length was installed, with a dump bailer, above a temporary mechanical plug. Groundwater in the hole, if any, was minimal during placement. After 13 days of curing, the grout mass was partially drilled out from both ends, leaving a seal 10 cm long. Water distribution in rock mass The distribution of water and fluid pressure in the rock mass near the seal appears to be complex. The generally dry condition of the borehole and rock mass suggests that the rock mass near the seal was unsaturated prior to the injection tests of this research. The injection tests presumably created an expanding saturated zone around the seal. However, anomalously high fluid pressures, consistently recorded by one of two piezometers installed near the seal, suggest connection to a source of significant natural hydraulic head. Comparable head variations for closely spaced piezometers have been reported by others (Sowers, 1976). [ ] Seal testing The seal was tested for 3.5 years, using steady constant-head, tracer travel-time, dye-injection, head-buildup and transient constant-head tests. These tests are fully described in Greer and Daemen, 1991. In the head-buildup test, transient flow occurs through the seal (and rock mass) to a closed, water-filled interval of the borehole adjacent to the seal. The test consists of monitoring the rise in head in the closed interval until steady conditions are approached. The test may be analyzed in terms of a one dimensional analytical model to obtain the effective hydraulic conductivity (Ks) and specific storage (Sss) of the seal. Analysis in terms of an axisymmetric three-dimensional numerical model may yield properties of both the seal and rock mass.

Jan 1, 1993

By T. J. Cox

The Red River is located within the Sangre De Cristo Mountain Range in northern New Mexico. An approximate 12-mile segment of the river from the town of Red River to the village of Questa (Questa) includes several hydrothermal alteration zones. Alteration ?scars? formed as a result of rapid chemical weathering and physical erosion of pyrite-rich hydrothermally altered rocks. Storm water runoff from the scar drainages contributes metal-laden, acidic water to the river and creates a visible plume of discolored water in the river. Acidic runoff reduces the river pH from 7 to the mid 3s for at least 2 hours during large storms and the low pH water continues downstream. Runoff from scar drainages increases some metal concentrations in the river from below chronic ecological criteria to as much as two orders of magnitude greater than acute ecological criteria, and results in acutely toxic conditions for macro-invertebrates. A general improvement in the river?s water quality occurs downstream due to dilution from tributary inflow from non-mineralized watersheds, the natural buffering capacity of the river, and storm water controls at the Chevron Mining Inc. Questa mine.

Jan 1, 2011

By James R. Jones

The concern for air pollution goes back centuries as will be seen from this quotation : "Strife and coal, it seems, have a hand-in-hand historical relationship. It was thought by some . . . in the Middle Ages that burning coal filled the air with dangerous poisons. So great was the concern that in 1306 King Edward I of England passed the first and unquestionably the harshest air pollution legislation edict that history will record. His proclamation declared the use of coal an act punishable by death." As related to coal mining, the air pollution problem is primarily that of its customer when using the fuel. Thus the major impact relates to markets for coal rather than the actual mining operation itself. In the vicinity of the coal mine the problem is principally related to fugitive dust from roads, transfer facilities, loading facilities, crushers, etc., and particulate matter (dust or fly ash) from thermal dryers and air cleaning tables.

Jan 1, 1981

By David B. Cook, James R. Bieke

I. INTRODUCTION Lawyers are a cautious breed; and it might be expected that a paper concerning a broad and complex area of legal regulation, prepared by lawyers for presentation to a predominantly non-legal audience, would commence with several caveats. Ours are the following: First, we have written a broad (and, in some instances, simplified) overview of the regulatory system and of some of the more important requirements generally applicable to uranium mine and mill wastes in the United States. Our paper is not a legal opinion setting forth all of the governmental requirements that would have to be met by any particular mine or mill. Similarly, we do not imply that the requirements we mention are necessarily applicable to each and every mine and mill in the country; depending on the facts, some of the requirements may be inapplicable or avoidable through exemption or variance proceedings. Compliance decisions and strategies for particular facilities must be considered on a site-specific basis. Second, we have undertaken merely to describe the regulatory framework and some of the generally "applicable" requirements which have been promulgated and proposed by administrative agencies. We express no opinion on whether, under the U.S. Constitution and statutes, the existing and proposed agency regulations are valid in general or as applied to any particular mining or milling operation. There are, of course, procedures for judicial review of agency action; and newly promulgated environmental regulations have frequently been challenged, with varying degrees of success, before the courts. For us even to begin to consider such issues would expand an already forbiddingly long paper to totally unmanageable proportions. Third, in light of the subject of this Conference we have limited our paper to the legal framework and requirements applicable to uranium mine and mill "wastes" -- primarily mine overburden and waste rock and mill tailings (including waste tailings solution). We treat operating discharges (e.g., emissions from yellowcake drying) only peripherally. Fourth and finally, it is important to note that this paper, at least in so far as it concerns the specific requirements of particular regulations, will likely soon be out of date in some important respects. Despite

Jan 1, 1980

By Victor L. Steven

10.1-SHAFT SINKING AND RAISING BY CONVENTIONAL METHODS VICTOR L. STEVEN 10.1.1-SHAFT SELECTION A great deal of thought and planning should go into the selection of a mine shaft to provide the most effective use over the greatest, period of time. It sometimes is better to make an additional capital expenditure in the initial shaft program to achieve the greatest overall shaft usage. It is possible to sink a shaft which could be used for mine development, and initial production and later convert it to a ventilation, service or escape shaft. It is important to consider a second shaft in the original planning to provide ventilation, an escapeway and services for overall efficient production and safety. A second shaft usually is required to comply with state and federal mining laws. The main items to consider in shaft selection are: 1. Installation from surface or underground. 2. Vertical or inclined (sec Sec. 10.1.4L for discussion of factors involved in choice). 3. Tonnage to be hoisted. 4. Size of mining equipment to be lowered through the shaft to the underground workings. 5. Amount of water to be handled in the mining operation. 6. Ventilation requirements. 7. Type of ground the shaft is to be sunk through. 8. Position of shaft with relation to ore body. After the main features of a shaft arc determined, such as vertical or inclined, shape, size, type of lining and support, compartments, and shaft. facilities (air, water, ventilation, pump lines, power cahles, etc.), then :I sinking plant and procedures should be planned. At this point it should be decided whether the company will sink the shaft with its own crews or award the job to a competent shaft contractor. Company sinking generally will cost less if competent shaft men are available. Money also will be saved in plant costs, such as for hoist, headframe, compressed air, buildings, etc., as these will be part of the permanent plant and can be used in the sinking operation. Labor costs generally will be lower than contractor construction rates. The advantage of having a contract, or is the speed of completion of the project. Contractors usually have headframes and hoists ready to move on the job immediately and have trained crews ready to complete the project in the shortest possible time. Next, the type of surface sinking plant (hoist, headframe, etc.) should be selected and a determination made whether it is to be temporary or permanent. In most cases it is less expensive to install the permanent, headframe and hoist, and use them for sinking. The type of actual sinking equipment to be used should be evaluated for the most efficient sinking cycle, e.g., mucking method, drilling method, and lining or support means. The proper cycling of drilling, mucking and lining is extremely important in obtaining the most favourable sinking costs. 10.1.2 - GROUND STABILIZATION AND WATER CONTROL Prior to the shaft-sinking operation an investigation should be made into ground-water conditions. If water is anticipated, then pregrouting of the ground

Jan 1, 1973