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By D. Buranek, J. G. Wulff, F. W. Gifford

I. INTRODUCTION During the past few years, the state of the art of disposal of solid and liquid waste products from uranium mining and milling operations has experienced some rather interesting and significant changes. Some of these changes have resulted from technological advances in the milling operations and in the engineering aspects of the actual design and construction of the disposal facilities. However, the most significant changes, at least those which have had the greatest impact on the comparative cost and economics of modern disposal systems, have been environmental requirements and regulatory controls, which have become increasingly stringent in recent years. Without delving into the details of the myriad of new laws and regulations which now control the uranium industry, it can be stated with reasonable confidence that one of the principal purposes of more rigid state and federal controls over this industry is the basic desire to promote safer, more reliable and longer term means of effectively controlling and disposing of the potentially objectionable waste products resulting from uranium mining and milling operations. It is difficult to take issue with these basic goals because their need was spawned by past lack of awareness and inattention on the parts of both government and industry to the long term environmental problems and potentially hazardous conditions which can be created by limited planning and ill-conceived or poorly controlled and executed disposal practices. On the other hand, despite the good intentions of those who developed and those who are administering the regulatory controls now in effect, the basic danger exists that overgeneralization of the approach to disposal problems and overrestriction of the potentially acceptable solutions for uranium mine and mill waste disposal could have a significant, adverse economic impact on the mining industry and the ultimate consumers of their products. The regulatory pendulum, which perhaps swung too far toward lack of controls over uranium mill waste disposal in the 1940's, could now well be headed too far in the opposite direction, that is toward over-regulation in the form of restrictions based on technically unrealistic or weakly supportable

Jan 1, 1980

By David A. Hedderly-Smith

Sealaska Corporation, the Alaska Native Regional Corporation of Southeastern Alaska, was organized under the laws of the state of Alaska pursuant to the federal Alaska Native Claims Settlement Act of 1971. Large areas of Sealaska's 630,000-acre (250,000 hectare [ha]) fee subsurface holdings are underlain by carbonate, and while much of this material is assumed to be common variety limestone and lesser dolomite, at least two areas in the Silurian Heceta Limestone contain high-quality limestone at tidewater. This limestone has many potential commercial uses. On northern Prince of Wales Island, about 90 miles (144 km) north-northwest of Ketchikan, resource development at Sealaska's Calder deposit has established a reserve of 13 mil- lion short tons (st) (12 million metric tonnes [mt]) of carbon- ate rock containing an average of greater than 98 percent CaC03, approximately 0.5 percent MgO, less than 0.5 per- cent Si02, and approximately 0.15 percent Fe203, with insolubles averaging less than 0.5 percent and an average Hunter Brightness of 92.07 for the -150 mesh grind (with a bright- ness cutoff of 90). The deposit is open at depth and to the south. Ultimate reserves will likely be several times those currently established. In 1998, Sealaska began mining the Calder deposit, with about 100,000 st (90,000 mt) produced, and also completed a marine terminal one half-mile (800 m) from the deposit. The terminal can load ships up to 680 feet (200 m) in length at a rate of 2,000 st (1,800 mt) per hour. 0 m) in length at a rate of 2,000 st (1,800 mt) per hour. On Dall Island at Breezy Bay, about 60 miles (100 km) southwest of Ketchikan, Sealaska has identified a limestone reserve in excess of 150 million st (135 million mt) that (based on 18 surface samples distributed evenly over about a section of land) contains about 98 percent CaC03, and less than 0.5 percent MgO, 0.5 percent Si02, and 0.15 percent Fe203. Brightness of the Breezy Bay carbonate is considerably less than at Calder. Five samples of brighter material from Breezy Bay returned brightness analyses in the high 80s and low 90s; average brightness is probably below the mid- 80s. Sealaska has also identified a deposit containing more than 19 million st (17 million mt) of similar rock near View Cove, a few miles south of Breezy Bay.

Jan 1, 2001

By Alden B. Carpenter

Brines (a solution containing more than 100,000 mg/l dissolved solids) in sedimentary basins are genetically related to evaporites in one of three ways: (1) dissolution of evaporite minerals, generally halite; (2) interstitial fluids in evaporites which are expelled from their source rock during compaction; (3) incongruent alteration of hydrous evaporite minerals (carnallite). Most brines of commercial interest originate as interstitial fluids in halite-rich rocks. These fluids are initially rich in magnesium, sulfate, and potassium but evolve in composition to an Na-Ca-Cl brine during their migration to their present host rock. A particularly important reaction in this process is: 2CaCO3 + Mgt+ = CaMg(CO3)2 + Cat (1a) This reaction reduces the magnesium content of the brine and ultimately converts it to a calcium-rich brine. The loss of sulfate occurs by two mechanisms. One of these is by direct precipitation as calcium sulfate (either as gypsum or anhydrite): Ca 2+ + SO4 2- = CaSO4 (2a) Additional sulfate may be lost from the brine by sulfate reduction: S042 + 1.33CH2 + 214- = H2S + 1.33CO2 + 1.33H20 (3a) The loss of carbonate species occurs by precipitation as carbonates: H2O + C02 + Cat = CaCO3 + 2H+ (4a) The net result of reactions 3 and 4 is: Ca 2+ + SO4 2- + 1.33CH2 = H2S + CaCO3 + 0.33H2O + 0.33CO2 (5a) The loss of potassium from brines is caused by the formation of potassium aluminosilicates: 2K+ + Al2Si2O5 (OH)4 + 4SiO2 = 2H + 2KAlSi2O8 + H2O (6a) The hydrogen ions produced by this reaction dissolve carbonates: 2H+ + CaCO3 = Ca 2+ + CO2 + H2O (7a) The net result of reactions 6 and 7 is: 2K+ + CaCO3 + Al2Si205 OH)4 + 4SiO2 = Ca2 + 2KA1Si3O8 + H2O + C02 (8a) A study of brines from central Mississippi has shown that the composition of these brines is related to the interaction of pore fluids from evaporites with minerals in the enclosing sedimentary rocks.

Jan 1, 1978

By D. W. Gentry

Introduction Until recently, there were many jobs for mineral engineering graduates. Traditionally, minerals-related career paths available to university graduates stem from rather standard offerings in the disciplines of mining, metallurgy, petroleum, geology, geophysical engineering, and mineral economics. Within each discipline, there exists many career opportunities. In fact, until fairly recently, the most difficult decision for most graduates was choosing among all the available alternatives. Unfortunately, this relative state of euphoria changed quickly and dramatically within the minerals sector. Within two or three years, graduates from minerals-related engineering disciplines found it exceedingly difficult - and then almost impossible - to secure employment in the minerals sector upon graduation. At the same time, there were dramatic reductions in production requirements and, ultimately, mine closings. This resulted in massive layoffs and often termination of professional personnel possessing significant and valuable industrial experience. Obviously, this worsened an already bad employment picture for those hoping to begin a career in minerals. Certainly, there is no universal agreement on what the future offers for the US minerals industry or the career opportunities associated with it. However, there are a number of considerations that deserve study by students, faculty, administrators, and industrial personnel before minerals-related career paths are chosen. Where Will the Jobs Be and How Many? Everyone is aware that the minerals industry is very cyclic. It is not surprising then that job opportunities for new graduates also fluctuate. Typically, there is a two-to-four year lag between industry's engineering manpower needs and the academic community's ability to produce this talent. Due to the response lag, there will necessarily be years when supply and demand do not match. Employment Projections It is extremely difficult to anticipate what the key career paths might be within the next few years. Considerable speculation remains with respect to whether activity in the US minerals sector over the short-, intermediate-, and long-term will improve, remain status quo, or further deteriorate. However, some information pertaining to the prospects for career opportunities in mining engineering was offered by Tingley and Ghose (ME, Sept. 1983, p. 1274) from a study performed by Fenvessy and Schwab. Results from this industry-sponsored survey suggest that the US will experience 700 more mining engineers than needed in 1985, 1300 in 1990, and 1800 in the year 2000. These conclusions were based on the assumptions that future graduate levels remain flat and demand increases 3.6% annually. Educators quickly argued that depressed job markets, coupled with university student mobility, would negate the assumption of flat mining graduate levels in the immediate future. Student Mobility Student mobility from one engineering discipline to another is high. This is perhaps best demonstrated by the tremendous growth in programs such as computer engineering and computer science at the expense of enrollments in the more traditional engineering programs. The ability and willingness of students to quickly change engineering disciplines in response to existing and perceived job market demands have created significant student, faculty, and resource dislocations within universities nationwide. This has resulted in problems for university faculty and administrators responsible for effective use of

Jan 2, 1985

By W. E. Cobb

The financial implications of environmental liabilities are becoming staggering, particularly for mining companies. The Environmental Protection Agency's (EPA) Record of Decision (ROD) list for Superfund mining sites has some pretty hefty financial commitments, such as the $10 million operable unit remedy for the Yak Tunnel at California Gulch (in Colorado) or the $40 to 60 million remediation at Bunker Hill (in Idaho). These costs are driven by the large volume and areal extent associated with historical waste management practices at mining sites. (It's difficult to have a low-cost remedy for a 20-square-mile site.) Unfortunately, there are other ways of generating environmental liabilities besides Superfund, such as non-compliance with Clean Air Act provisions, stormwater permits, or National Pollutant Discharge Elimination System (NPDES) permits. In the last 2 years, EPA has increased enforcement in its regulatory programs, which has resulted in increased fines and criminal actions. Leaders in other industries (primarily the chemical and manufacturing industries) are viewing environmental liabilities as a strong motivator to select engineering approaches carefully based on technical, legal, economic, and philosophical perspectives. Industry interest in waste minimization programs comes from the concept that it is economically advantageous not to create waste in the first place. Waste minimization programs typically balance increased front-end capital investment with the minimization of potential long-term environmental liabilities as a way to achieve the lowest total-or life-cycle-cost. A side benefit of waste minimization programs is that they tend to improve overall regulatory compliance. Waste management practices are also becoming philosophical, as evidenced by a May 11, 1992, Business Week article: Companies that incorporate innovative thinking on sustainable development will do more than clean up or prevent their own pollution. They'll make and distribute a product more efficiently than ever, worry about its lifelong environmental impact, plan for recycling, and get their suppliers on the bandwagon. This type of approach will require a new mind-set, new tools, and adjustments throughout the organization. A key aspect of waste minimization or life-cycle waste management programs is the incorporation of environmental costs into economic analyses that compare various waste management alternatives. However, the incorporation of environmental costs into an economic analysis can be more difficult than initially thought. As noted above, new tools-or even a new toolbag-are needed, such as new accounting methods that track and assign the true lifetime environmental costs of a particular product. One piece of the new toolbag is life-cycle cost analysis, which is discussed in this paper. To simplify the discussion, this paper looks at the total costs associated with a given mining site. The concepts presented in this paper, however, can be applied to any mine waste management decision. The ultimate extension of life-cycle costing, as presented in the Business Week article, is to include the potential environmental costs throughout the life of a metal. At this point, I am perplexed on why, let alone how, to include the potential downstream environmental costs (i.e., the costs associated with smelting, manufacturing, use, and recycle or disposal) throughout the life cycle of mined copper, lead, or zinc.

Jan 1, 1992

By John A. Kaufman, Terry D. Vandell

Recent mining regulations have increased the need for detailed engineering and environmental information pertaining to the hydrogeologic environment in mining areas. The cost of obtaining such hydrogeologic information has also increased. After mineralization of potentially economic grade ore and tonnage has been identified, and after completing preliminary mining feasibility studies (suggesting that development is merited), a second exploration phase can incorporate a simple, cost-effective baseline hydrogeologic data reconnaissance program that is limited but useful. Using exploration drill holes only, this program can provide a direct indication of where ground water is encountered. Also, reasonable estimates can be made on the relative production of water from various water-bearing strata, a composite potentiometric surface map can be prepared, and the general chemistry of the ground water can be characterized. This article outlines what typical baseline hydrogeologic data are needed and describes how it can be easily obtained during exploration drilling. Introduction Major engineering and environmental ground water parameters can be identified early in a project by using, drill hole data obtained from a mineral exploration program, designed by the mine geologist and hydrogeologist. Most importantly, data obtained in the exploration drilling program can be used to design a logical, defensible, and cost-effective baseline ground water data reconnaissance and monitor well program, which would be required for mine design and ultimately for mine permit approval. Drilling Where possible, exploration drill holes should be drilled with conventional rotary drilling equipment using air or, if necessary, water as the drilling fluid. Rotary air or water drilling generally allow good delineation of the first major saturated or water-bearing zone. If mud must be used during drilling, distinct hydrogeologic zones can still be defined, although the task may be considerably more difficult. Water Levels Static water levels can be obtained from selected exploration drill holes at specific depths while additional boreholes are drilled. Such measurements should not, however, be conducted until the hole has been developed with air (no water) and has stabilized at each of the depths to be measured. Such water-level information can be particularly useful when more than one aquifer exists. Where mine development is eminent, the hydrogeologist should install 50-100 mm (2-4 in.) diameter PVC casing in select drill holes. The PVC should be slotted, gravel packed, and properly sealed so additional hydrogeologic data can be collected for specific hydrogeologic zones at a later date. Such data could include water level information, permeability estimates based on injection and /or pumping tests, and water quality data. Where casing is not installed in drill holes, or when the drill holes are abandoned, the holes should be properly grouted and back-filled to the ground surface to prevent potential future ground water contamination. Monitoring Air Lift Water Production Estimates of the relative production from various water-bearing strata can be obtained from selected exploration drill holes. If ground water is encountered, its air lift (ground water flow) production can be measured at distinct intervals using a surface water flow measuring device, such as a portable Parshall

Jan 2, 1983

By W. A. Crosby, A. Bauer

The recent increase in size of open-pit mining operations has resulted in major improvements in efficiency that are beneficial to the mining industry. Higher bench heights, larger diameter blast-holes, and more powerful explosives have played an important part in reducing mining costs. These measures have, however, resulted in increased energy concentrations in the blast area, which can result in severe backbreak problems for final pit walls. If backbreak is not controlled, a decrease in the overall pit slope angle will ultimately be necessary, and consequences, such as decreased recoverable ore reserves and increased waste-to-ore ratios, will result. Greater amounts of loose face rock will be produced, and planned safety berms will be less effective or nonexistent. Hazardous working conditions can also result. Remedial measures, such as scaling large areas and using wire mesh or other artificial support, are very expensive and difficult to implement. There must be a trade-off between the money saved by using larger blasts, and the time spent to maintain pit wall quality. The best approach is to control blasting effects so the inherent strength of pit walls is not destroyed. Presplit and trim blasting are two control blasting techniques used to accomplish this. Both techniques are designed to create a low explosive energy concentration per unit of wall area at the perimeter of the pit. Energy concentration of the main production blast must also be controlled to avoid damaging the final pit wall. A low-energy concentration at the final wall can be obtained by decoupling charges, decking charges, using less powerful explosives, decreasing blasthole diameters, and changing the burden and spacing. Rock properties-fragmentation and the final state of the pit wall-influence the success of a blast. The most important properties to consider are: in situ rock strength, and the nature, frequency, and orientation of structural features. Since these variables cannot be controlled, they must be evaluated by suitable field tests, and then a control blast can be designed using controllable variables, such as spacing, burden, and hole diameter, so the blast will be successful for appropriate rock conditions. One main thrust of this article is to show how control blasts can be successfully designed for larger hole sizes, up to 381-mm-diam. While efforts to produce a good control blast in shovel/truck pits have been ongoing for some time, recent use of a modified presplit technique in dragline operations has helped reduce water problems, improve safety, and increase dragline productivity in a number of surface coal mines. This article presents examples on designing and using presplit and trim blasting for large hole operations in both shovel/truck and dragline mines. Presplit Blasting As introduction to presplit blasting, Fig. 1 illustrates a typical presplit blast layout using 102 mm-diam presplit holes for 381¬mm-diam production holes. For this type of blast, presplit holes would normally be drilled first, ahead of main production holes. The choice can then be made between loading and firing the presplit line or infilling the main blast. In the latter case, the presplit line would be fired instantaneously 100-150 milliseconds before the main blast. As shown in the figure, the presplit line is formed ahead of the main

Jan 2, 1982

By Michael T. Ballard

The key to productive computer usage is system software integration with full compatibility among individual components. The majority of software developers in our industry are still in the process of perfecting individual components. As a result, exchange and validation of data between disciplines becomes an arduous process. We at Morrison Knudsen Engineers, Inc. (MKE) have integrated an entire system, allowing exploration, geology, modeling, mine planning, and operation modules to work as a single, complete system with direct exchange and validation of data. The importance of integration to both developers and users of today's mining soft- ware is becoming increasingly apparent as we, an industry, move toward growth and enhancement of the technology of ore reserve acquisition, evaluation, and management. It is also important to understand the present state of integration and where it is leading US. Not long ago, another industry passed through the same stages we are now setting up for the mining industry. The stereo industry began with a few individual components, the record player, tape recorder, and radio. People were able to play records, listen to the radio and record conversations and music only through microphones on their tape recorders. Soon, one enterprising individual discovered he could tape sound directly from the radio or record player by capturing audio output from the speaker wires. Interfaces and patch cords were developed, providing a means to transfer audio information between components with not only little loss of sound quality, but also the elimination of back- ground noise. The industry took another step forward with the introduction of "integrated" stereo, where the components were packaged in one cabinet with shared resources. This reduced overall cost which paved the way for better quality. Today's stereos are now so well integrated that many of the components share data and resources from the same manufacturer, designed to work together for almost perfect data transfer between any of the components. The result is better sound, convenience, and reduced cost. This is the ultimate goal for integration. Any system development passes through these same steps on the road to total interaction. Initially, the components are designed, tested, and marketed individually, often by separate companies. As the users become familiar with the components, the need arises for an automatic interface between the components. This is necessary to cut down on error during data transfer and to eliminate the redundant work required to enter results from one component into the next component in the system. As a result, interfaces must then be developed to aid in this data transfer. After data transfer is accomplished, the next step is resource sharing. Resource sharing is careful design and engineering by either cooperative, individual manufacturers or one firm taking on the task to produce a single package product. Design and resource sharing leads to lower overall cost, while still providing component capability with options to grow as the user grows. The mining software industry has been working to develop some very high quality component software and has indeed accomplished this task in every aspect of the mining industry. However, realizing the importance of integration, some companies with limited individual resources have started working together to interface their individual components. Other firms have undertaken the project of developing complete, integrated systems.

Jan 1, 1987

By A. Haile, A. Poole

During 2001-2002, the Granny Smith Mine of Western Australia undertook the development of a planned circular failure. Excavation took place in an interim pit wall that intersected the transported sediments, lake clays, and palaeo channel sands that overlay the deposit. The purpose of the study was to refine feasibility level slope parameters used for the initial design of the Wallaby pit, with the intent of increasing the overall slope when excavating the ultimate pit limits. During excavation, the geological and hydrogeological characteristics were studied in detail and data was collected using area-specific instrumentation and monitoring. Failure occurred in mid February 2002, at which time a detailed survey of the excavated slope using laser technology was conducted to provide geometric information suitable for back analysis of the material shear strength. The slope failure and associated information was analysed by geotechnical consultants. The revised shear strength and hydrogeological parameters were then applied to design the final pit walls. The slope redesign provided a cost saving of approximately seven hundred thousand US dollars. Design, slope formation, back analysis, the geology of the clay units, the hydrological environment, planned infrastructure, management and safety issues, and the on-going slope performance of the final slope are discussed.

Jan 1, 2004

By J. W. Leonard

Introduction Gravity-based coal/mineral preparation plants must be operated to yield products of a quality satisfactory to existing markets. However, because of historic control problems, recovery is frequently sacrificed to achieve quality, resulting in losses of valuable product. Plant control has ranged from intuitive in-plant spot quality judgments made by various operators to gross methods involving such devices as density gauges, floats, grabsample on-the-spot sink/float tests, etc. Millions of tons of quality coal/minerals are lost to refuse each year, because of the historic lack of a continuous and dependable control system. The result is serious economic, as well as environmental, losses. While the economic losses are obvious, the less obvious environmental ones, in the case of coal, result from the misplacement of more ignitable quality fuel to refuse banks. Until recently, a control system capable of supplying a market-grade coal/mineral product, while continuously minimizing product losses to refuse, has been technologically unavailable. Coal/Mineral Washing Tracer System Performance Monitor (CMWTSPM) The first Coal/Mineral Washing Tracer System Performance Monitor (CMWTSPM) originated in the late 1970s as an innovative idea and ultimately resulted in Leonard Patent No. 4,345,994 issued on August 24, 1982. In a classic example of successful university basic research, the patent was followed by two US Department of Energy (DOE) merit awards for fundamental optimization research at the University of Kentucky, a number of publications, a successful applied research program (underwritten by SENTREX at Irvine, CA) and field tests at the Rochester and Pittsburgh Coal Company Keystone Mine near Indiana, PA. The system is now operational and is marketed by SENTREX as the SENTREX Performance Monitor. A diagram of the system is shown in Fig. 1. This system relies on metal detection as the basic tracing medium. Since it is capable of tracing particles as fine as 0.19-in., the system is ideal for coarse coal/ mineral cleaning processes. Basically, the CMWTSPM records the distribution of synthetic coal/mineral particles, of known size and specific gravity, between product and refuse output streams of coal/mineral washing, gravity-concentrating processes. For any given particle size, surface area and shape are constant. The particles are [ ] designed to find their way into the center of their respective inprocess specific gravity/size streamlines. The system will yield a complete partition curve to show the percentage of misplacement for any selected size/specific-gravity particle fraction, thereby providing a point-to-point insight into process failure. When used to evaluate operational processes, this full partition curve is far more diagnostic than the simplifying, derivative, single index number known as the Ep value. The Ep value is commonly used for convenience to represent symmetrical partition curves, but, if the actual partition curve is non-symmetrical, use of the Ep value can result in serious diagnostic error. While the Ep is useful in design and for rapid performance evaluation and process adjustment, the full partition curve is by far the more diagnostic. It will provide insight into the behavior of particulate flow, through non-symmetry, as a means to highlight potential systemic failures. Performance of the process is measured and improved by determining the extent of particle misplacement and by making timely adjustments to the unit to reduce the misplacement. Used in this manner, the CMWTSPM is not only a long needed processing control monitor, but it is also a very powerful diagnostic tool which, in the hands of skilled operators, has

Jan 1, 1993

By G. R. Laurer, J. W. Neton, C. Pomroy, N. Cohen, M. Eisenbud

INTRODUCTION In uranium miners, a causal relationship between exposure of the lung to high levels of alpha radiation and the consequential induction of bronchial carcinoma is now well-established (Wagoner, et al., 1965). Quantitation of this cause-effect association presents considerable difficulty, however, as the result of various confounding factors such as the contributing effects of smoking and/or arsenic exposure (Archer, Gillam, and Wagoner, 1976), normal metabolic variability and, perhaps most important of all, the absence of accurate radon/radon daughter exposure data. Estimates of cumulative lung exposures suffer from inaccuracy due to critical factors such as 1) unknown ventilation rates; 2) difficulties in obtaining representative mine air samples; 3) the variable character and magnitudes of the exposure sources; and 4) specific differences in metabolism and job histories that exist among miners. It was with considerable enthusiasm, therefore, that investigators readily embraced the idea of using skeletal burdens of the relatively longlived descendants of radon and its short-lived daughters, i.e., 210 Pb and 210Po (Archer, et al., 1968), to provide a measure of cumulative past exposure wherein the body acts as an integrating dosimeter. Early studies considered excretion rates of 210Pb and 210Po in urine, blood and hair (Sultzer and Hursh, 1954; Black, et al., 1968; Jaworowski, 1965; Bell and Gilliland, 1964) as well as their concentrations in bone samples, if and when they became available (Blanchard, Archer, and Saccomanno, 1969). These methods were ultimately complemented by in vivo, whole-body counting techniques using low background, NaI-Csl(T1) detectors (Eisenbud, et al., 1969). By this latter method, it would be possible to measure the skeletal burden of 210Pb directly, thereby avoiding the many difficulties associated with the relation between nuclide excretion rates and body burden. The rationale was simple; by knowing the skeletal burden of 210Pb it should be possible, by the judicious use of an applicable metabolic model, to estimate the cumulative lung exposure to alpha-emitting radon daughters (Fig. 1). Various models that were consequently constructed, however (Fisher, 1969), all suffered from the same limitation, i.e., the assumption that the source of 210 Pb in the skeleton and the source of lung exposure to the short-lived radon daughters were simply related and, if not, were at [ ] least not too variable (Raabe, 1970). By now, the difficulty with this assumption has been recognized and more sophisticated modeling techniques have been attempted which presumably present a more realistic picture of exposure conditions (Holtzman and Rundo, 1981). There is, nevertheless, the pessimistic view that due to the variety and variability (in time as well as in character) of sources that may exist for the origin of 210Pb in bone, it may [never] be possible to estimate past lung exposures accurately by an indirect, predictive model. For example, the origin of 210 Pb in the skeleton may include such sources as: 1) the decay of short-lived 222Rn daughters, inhaled and ingested; 2) the decay of 222Rn dissolved in body fluids and tissues; 3) directly from 210 Pb present in air, inhaled and ingested; and 4) 210Pb from non-occupational sources, e.g., food, water, smoking and endogenous 226Ra. The contribution of each of these sources may vary depending upon specific site conditions such as ventilation and mining operation. On the other hand, empirical evidence seems to indicate a trend reflecting the simplest of direct

Jan 1, 1981

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

By E. P. Fahrenthold, K. E. Gray

Introduction Brine reservoirs along the Texas and Louisiana Gulf coasts are currently under study to investigate the technical and economic viability of production wells designed to extract methane gas in solution. The effect of well pressure drawdown on reservoir rock and associated overburden material is important to well performance predictions and the analysis of potential subsidence problems. Consequently a series of compaction tests have been conducted on sandstone and shale cored from a test well drilled near Hitchcock, Texas, in order to determine the variation of mechanical properties to be expected over the producing life of the formation. Previous Work The referenced tests are part of an extended study of geopressured-geothermal reservoir behavior, including compaction (Jogi, Gray, Ashman, and Thompson, 1981 and Jogi, Kalra, Gray, and Thompson, 1981) and creep (Thompson, Jogi, Gray, and Richardson, 1981) tests of reservoir rock and related reservoir performance studies (Rago, Okhuma, Seperhnoori, and Thompson, 1985). Recent work has considered the nonlinear constitutive behavior of sandstone (Fahrenthold and Gray, 1985 and 19861, while additional experimental work is reported here. Geological structure and composition of the Hitchcock field from which the core samples were obtained is discussed by Light (1985). Experimental Procedure Compaction tests were conducted on cylindrical cores, nominally 2.75 inches in length and 2 inches in diameter, with independent control of the axial, confining, and pore pressures. Axial and radial deformation of the samples were measured in tests over a range of effective stress states expected to develop over the reservoir lifetime. The term effective stress as used here refers to the algebraic difference of the total stress and pore pressure, with compression taken as positive. Three types of compaction tests were conducted on the sandstone. In the uniaxial tests, axial total stress and radial strain were held constant as the pore pressure was reduced, similating one dimensional compaction which may occur during reservoir depletion. In the triaxial tests, radial total stress and pore pressure were held constant as the axial total stress was increased, duplicating conventional rock compaction test procedures. In the pore pressure drawdown tests, axial and radial total stresses were held constant while the pore pressure was reduced, in order to investigate differences in sample response to mechanical and pore pressure loading. Compaction tests were also conducted on shale material obtained from the reservoir overburden. The field coring process provided few consolidated shale sections sufficiently large to allow preparation of samples with the dimensions described above. With additional attrition during sample preparation of this brittle material, ultimate yield of the coring effort was limited. The few available samples were tested in uniaxial and triaxial tests similar to those described above. The use of displacement transducers to measure sample deformation avoided the difficulty of strain gauge installation on this material. Test Results The sandstone and shale both demonstrated nonlinear stress-strain behavior, with hysteresis and residual strain present upon unloading. Polynomial interpolation of the stress-strain curves was used in the calculation of incremental values for the loading and unloading moduli. Figures 1 and 2 show example axial effective stress-axial strain curves for triaxial tests conducted on sandstone and shale samples. The figures plot incremental stress and strain variables, defined as deviations from the state of prestress to which the sample was subjected. The state of prestress was varied between tests in order to investigate material behavior over the entire stress and strain space of interest. Tables 1 and 2 give the incremental moduli (Young's modulus and Poisson's ratio) as a

Jan 1, 1986

By J. A. Siekmeier

The thickness of the laminated beam modeled in this study is based on bridging potential and TDR measurements. The laminated beam stiffness is based on the effective stiffness of an equivalent transformed section. The maximum deflection of this near-surface laminated beam is compared with the maximum measured surface subsidence. The spreadsheet provides a rational technique to graphically display the character of the overburden and potentially allows prediction of a subsidence factor based on boring logs and the proposed mine plan. The U.S. Department of the Interior, Bureau of Mines (Bureau), utilizes a commercially available computer spreadsheet program in combination with a modified version of Bieniawski's Rock Mass Rating (RMR) system to characterize the geology overlying high extraction coal mines. The spreadsheet program calculates a RMR for each bed based on the lithology, thickness, and engineering properties, which are determined from boring logs and laboratory tests. Once the RMR is calculated the spreadsheet has two major functions. First, the spreadsheet calculates values used to graphically represent the character of each individual bed in the overburden. It calculates an in situ deformation modulus using an empirical rela¬tionship between the modulus and the RMR and then uses this effective modulus along with additional parameters that are based on mine geometry to calculate the bending stiffness and bridging potential for each bed. Second, the spreadsheet creates a transformed section that allows groups of beds to be modeled as an elastic beam. It then calculates a bridging potential for the entire rock mass. The bridging potential of the rock mass is a single value that takes into account both the geology of the overburden and the geometry of the mine. It seems reasonable that there would be a strong correlation between the bridging potential, the overburden response to high extraction mining, and the subsidence ratio. The results presented in this paper focus on the relationship between the bridging potential of individual beds and the deformation measured using Time Domain Reflectometry (TOR). Subsurface displacements measured over seven high extraction coal mines in southern Illinois were correlated with the bed stiffness and bridging potential profiles. Based on these correlations it was evident that the process of caving and subsidence over these high extraction mines involved momentary bridging by individual beds or groups of beds and the ultimate development of a laminated beam near the surface.

Jan 1, 1986

By W. Robert Power

Stone is considered by many to be the premium material for beauty and durability in all kinds of construction. Its use dates to the dawn of civilization and only buildings made of stone have survived from ancient times. The use of stone relative to other materials declined from the beginning of the present century reaching a low point in the third quarter, but there has been a resurgence in the use of stone in construction during the past decade, due probably, to new techniques for producing stone, a greater recognition of its beauty and durability, and greater affluence among consumers. The aesthetic appeal of stone is attested to by the many attempts to simulate it in synthetic materials. Today, natural stone is considered by many to be the ultimate in aesthetic appeal, durability, and ease of maintenance. Consumption of dimension stone has increased dramatically during the past decade. The US Bureau of Mines reported that the total value of US consumption in 1978 was $1 13 million and that total imports were valued at $65 million in 1979 (Anon., 1980). In 1989 consumption of domestic dimension stone was valued at $208.3 million and imports at $525 million. This dramatic increase occurred after demand had decreased at an average rate of 2.5% in the two decades prior to 1978 (Anon., 1980). More than half of the imports come from Italy, but production of dimension stone has increased worldwide spectacularly. Meaningful statistics on dimension stone are hard to acquire. The finished piece of stone is a highly value-added product, but stone may be sold as rough blocks, sawed slabs, or finished product. The same piece of stone may therefore be sold and reported three or more times. Furthermore, although the quantity of stone is most commonly reported in tons, stone cladding is more often sold by the surface area and at the same price for different thicknesses. By any measure, however, the production and consumption of dimension stone has increased dramatically during the past decade. A significant indication of the resurgence of stone was the introduction of two new trade publications during the 1980s. Stone World (Tradelink Publishing Co., Inc., Oradell, NJ), introduced in 1984 is published monthly and Dimensional Stone (Dimensional Stone Institute, Inc., Woodland Hills, CA), introduced in 1985 is published monthly. Both magazines provide useful information on new technology, market trends, producing areas and companies, and other information not previously available.

Jan 1, 1994

By Vincente M. Jr. Lim, Brain W. Lawrence

Underground mining operating costs consist pri¬marily of direct costs such as labor, supervision, ma¬terials, and supplies, plus indirect costs. The material and supply costs vary substantially between mines but commonly range from 35 to 60% of the total direct costs. As part of the preparation of accurate operating cost estimates for a project, information must be as¬sembled to select the material and supply items and to determine the quantities of these items that will be required. Cost data then can be incorporated with the estimated material and supply requirements to determine the total material and supply costs for the project. Estimated quantities will vary considerably between mining projects depending upon factors such as the type of material being mined, the mining method, the equip¬ment selected, drilling and blasting techniques, mucking and haulage methods, ground support requirements, number of men employed, etc. Valuable information often can be obtained from suppliers and manufacturers, including data on current operating practices for similar projects, recommenda¬tions for the specific project, and preliminary cost quota¬tions. However, quotations obtained for estimates are generally higher than the prices paid at the time of purchase. Mines operating in the area, or those mining under similar conditions, also can provide essential design information, but it often is difficult to obtain cost information from the operators. Estimated costs for the most common major items of underground materials and supplies are presented in the accompanying tables. They represent current costs or cost ranges and are intended to be used only as a guide when preparing budgetary estimates for materials and supplies. Cost data were obtained mainly from suppliers and manufacturers, with up to three quotations being ob¬tained for each item whenever possible. In most cases, as noted in the table footnotes, prices were based on quantities representing normal requirements for me¬dium-size mine operations. Every attempt was made to assure accuracy. However, the cost for most items, especially those with multiple sources, can deviate sig¬nificantly for various reasons such as geographical loca¬tion, current market conditions, quantities ordered, size of the customer (which could influence the ultimate quantities purchased), and the ability of the purchasing department. It is also for these reasons that the actual amount individual mines are now paying for specific items may vary from the estimated cost in the tables. The cost for specific sizes of items that do not appear in the tables often can be factored from the data pre¬sented. Costs for many items can be converted easily to a unit that sometimes is more basic, but not as com¬monly used, such as from a length to a weight unit. Other items not listed, such as repair parts for equip¬ment, will involve contact with suppliers to determine specific recommended requirements and to obtain costs for these items as related to each individual project. Costs are expressed in terms of US dollars as of 1978 unless otherwise specified. All measurements, sizes, and weights are expressed in conventional units (i.e. feet, inches, pounds) ; applicable metric conversion fac¬tors follow each table. Costs are represented for drill steel and bits; mining explosives; blasting caps; blasting supplies; ground sup¬port items; ventilation equipment; utility pipe, utility hose, and accessories; mine track and accessories; tires for underground equipment; lubricants and fuels; and miscellaneous supplies.

Jan 1, 1982

By William W. Wellborn

Abrasives are substances both natural and synthetic that are used to grind, polish, abrade, scour, clean, or otherwise remove solid material usually by rubbing action (as in a grinding wheel), but also by impact (pressure blasting). Facts about the discovery, early use, adaptation, and development of abrasives prior to the 20th century are few. From the sparse, earliest information that is available, it would appear that the science of abrasives began when, to fashion his tools, man began to select certain kinds of rocks in preference to others because of their peculiar properties. There is evidence that stones were sawed by some sort of grinding machine as early as 4000 BC in Egypt. The real expansion of abrasives use coincided with the beginning of metallurgy in the Middle East around 2000 BC. The earliest confirmed use of abrasives for grinding metal comes from a steel dagger found along with a sharpening stone which was dated at about 1500 BC. Natural and manufactured abrasive substances play an important role in fashioning and finishing numerous products with a wide variety of end-use applications. The mining and preparation of crude, natural abrasives for their markets and the manufacture of synthetic and other abrasive materials contribute significantly to the US economy (Table 1). In addition, just as sawing to specific dimensions enhanced the value of the Egyptian stone, the use of abrasives in fabrication of modem products, and the addition of abrasives to others, add value. GENERAL CONSIDERATIONS The most important physical properties of materials that qualify as abrasives are hardness, toughness (or rigidity), grain shape and size, character of fracture (or cleavage), and purity (or uniformity). To fabricate bonded abrasive products such as grinding wheels, additional considerations include stability under high heat and bonding characteristics of grain surfaces. The economic factors of cost and availability are always important. No one singular property is paramount for any use. For some applications extreme hardness and toughness are needed, as in diamonds for drill bits. For other uses, factors of greatest importance are hardness and ability to break down slowly and, in the process, develop fresh cutting edges when grains become worn. In garnet sandpaper, for example, highly cleavable, friable, or extremely tough grains are not desirable. For other uses, extreme hardness may be objectionable. Abrasives for dentifrices and for glass-cleaning soaps are cases in point. For the most efficient uses in more critical applications, different types of abrasives are rarely completely interchangeable. While crushed quartz and garnet are both used in sandpaper, the two abrasives are not at all interchangeable in their applications. In the final analysis, the choice of a high-grade abrasive depends upon the quality and quantity of work performed by the abrasive per unit of cost. Initial cost of an artificial abrasive may be much greater than that of a natural abrasive, but the artificial mineral may do so much better work and do it so much faster that the ultimate cost is less. It is for this reason that artificial abrasives have largely replaced natural abrasives.

Jan 1, 1994

Throughout the 1970s, the US Bureau of Mines rode an ebb tide in Washington political circles. In rapid fire succession, the agency lost its mine health and safety enforcement responsibility (1973), a large portion of its energy research responsibility (1975), and then its fuel statistics functions plus a sizable portion of its coal R&D (1977). By 1978, johnny-come-latelies in the Energy, Labor, and Interior Departments had siphoned away from the Bureau some $125 million in program responsibilities as well as 900 of its employees. The bruising decade came to a close with a Department of Natural Resources proposal that, in its original form, would have scattered the Bureau's responsibilities to the winds. "There is no doubt in my wind," said Lindsay Norman, acting director of the Bureau of Mines (see box), "that were the Bureau to continue down the path it was on, the next step was closing the doors and going out of business." The doors are still open, but erosion of responsibilities had a devastating effect on the resources available to fulfill the Bureau's mission. Devastating, too, in terms of internal morale, considering the Bureau had six directors in a 10-year span. "When the troops are in the trenches and are being shot at," Norman said, "they need a leader-hopefully, a permanent one- that can get them advancing. This is something the Bureau did not consistently have in the 1970s." In September 1978, Roger Markle climbed aboard as director of the Bureau to rejuvenate the 70-year-old agency. He resigned only six months later* but the program Markle set in motion is being carried through by Lindsay Norman. Since Norman's appointment as acting director, he has spent considerable time on the road explaining his agency's redefined goals. At one such stop in Duluth, Norman and his two deputy directors, Charles Kenahan and Mike Kaas, answered the following questions. Lindsay, you were the principal architect in the Bureau's 1979 reorganization. How about a glimpse of what that entailed? Toward the end of the 1970s, after almost a decade of steady dismemberment, we reached what we felt was a critical point in the Bureau's history-we had the choice of redefining and strengthening our organization to meet the mineral challenges ahead, or continuing down the path toward the agency's eventual demise. We felt the Bureau's mission too important to abandon, so in 1978 with Roger Markle as director we initiated a six-month, no-holds-barred intro spection of just what the Bureau should be doing, why, and toward what ultimate national needs. And what did this self-reflection reveal? That there is, indeed, a unique federal role for the Bureau of Mines-that of a lead agency for influencing mineral decisions and pioneering tomorrow's technologic advances. Don't confuse us with policymakers, we're not. But we are policy analysts; we're scientists, engineers, and mineral professionals who can offer a sorely needed source of objective, accurate information to decision makers in both the public and private sectors. I feel there is no better brain trust in government than the Bureau when it comes to mineral expertise-of course, I'm biased. Do I understand you to say the Bureau of Mines is an advocate agency for the minerals industry? We plead no parochial interests, nor ought we, but from our viewpoint a sound minerals industry is extremely important to the broader national goals of security and a healthy economy. US dependence on foreign sources for critical minerals, the increasing noncompetitiveness of some US mineral sectors, and the array of land use issues surrounding mineral-rich lands all have much more serious implications than many otherwise informed people seem to recognize. In your estimation, how vulnerable is the US to foreign mineral suppliers? Let me make a distinction between dependency and vulnerability. The

Jan 4, 1980

By J. E. Moyle

Introduction In October 1983, Rio Algom Ltd. decided to develop the East Kemptville tin deposit. This decision was the result of intensive geological and metallurgical evaluation work, as well as careful studies of environmental and marketing aspects. The property is located about 54 km (34 miles) northeast of Yarmouth, Nova Scotia, near the village of East Kemptville. Geology The geological setting is a granitic intrusive at the contact with metasediment. The tin mineralization is in the form of cassiterite (SnO2), which is about 80% tin. Sulfides are associated with the deposit containing silver, copper, and zinc, which are recoverable as byproducts. The mineralization is associated with quartz veins and greisen alteration. In the center of the deposit, the zones are so close together they have the appearance of a massive or porphyry type deposit. Ore Reserves The mine plan is based on 56 Mt (62 million st) at 0.165% tin. Up to the ninth year, mill feed will average about 0.2% and lower grade material will be stockpiled. In later years, the average grade will be 0.16% tin. A 0.08% cutoff grade is planned for the entire 17 years of production. East Kemptville is a relatively low grade tin deposit, and only a combination of favorable circumstances has made it economic. • The ore body is situated at, or very close to the surface, permitting mining by open-pit methods at a very low stripping ratio. • The tin occurs in relatively large crystals that allow for good recovery by gravity techniques. • It is ideally located in a well-settled area, with all essential infrastructural requirements already in place. • The positive and cooperative response from various governmental agencies involved are encouraging. Mine Plan The mine plan incorporates an initial pit, followed by three expansions to the ultimate pit. Mining will be at 14 kt/d (15,000 stpd) and will include 9 kt (10,000 st) of ore plus 5 kt (5,500 st) of waste and low grade. After five years, the mining rate will increase to 21 kt/d (27,500 stpd). It is planned to drill vertical 165-mm (6.5-in.) blast holes on a 6-x 6-m (20- x 20-ft) pattern. Bulk slurry explosives will be used to break the rock. Loading will be with hydraulic shovels into 70 t (77 st) haulage trucks. The trucks will deliver the ore to a 13 x 20 m (42 x 65 ft) gyratory crusher, where it will be crushed to 150 mm (6 in.). Secondary and tertiary cone crushers will be used to further crush the ore to 13 mm (0.5 in.). Process Design Tin extraction technology is practically unknown in North America, but Rio Tinto Zinc, which has a 53% ownership in Rio Algom, is involved in tin production in the UK. Rio Algom has used this experience and expertise in metallurgical evaluation work and in tin-related aspects of plant design. Mineralogical and metallurgical test work have been performed, both on laboratory and pilot plant scales. Mineralogical work was done largely in the UK, the laboratory test program was carried out at the Lakefield research facility in Ontario, and pilot plant work was conducted in Australia. This program has resulted in significant improvements in recovery and concentrate grades, over best projections that could be made before this work was carried out. Concentrating Process The concentrator is being designed to process about 9 kt/d (10,000 stpd). The first step is to reduce the size of the ore to -700 µm (-24 mesh) in a rod mill/ball mill grinding circuit. The ball mill will operate in closed circuit with high capacity banana-type screens. The screened product will be deslimed to 30 µm (560 mesh) with cyclones, then passed over the pri-

Jan 4, 1984

By M. M. Botz, F. W. DeVries

Discussion by F.W. DeVries Comments on natural cyanide attenuation. The paper by Botz and Mudder in the November 2000 issue of Minerals and Metallurgical Processing had one disappointing aspect. Table 1 (p. 229) failed to include a significant species - neutral copper cyanide (CuCN). This compound has been most trouble-some in many plant-scale cyanide cleanup processes, includ- ing acidification-volatilization- reneutralization (AVR), hy- drogen peroxide cleanup and air/SO2 cleanup. It has also plagued many pilot efforts to use ion-exchange (IX) resin technology. This author elucidated the concept that, under some conditions, copper cyanide failed to precipitate out of treated effluents and appeared to exist as a filterable colloid that reported as copper and cyanide, with the latter measured as WAD (weak-acid-dissociable) cyanide in that analytical method. The concentrations measured were very near the levels of 0.2 ppm, often the regulatory limit for WAD cyanide. The copper concentration, using the 1 : 1 ratio for this com- pound, would then be more than 0.4 ppm, which could also be unacceptable. We discovered that it was unsafe to assume that the Cu(CN),- would surrender both cyanide ions simultaneously and allow cupric copper to be removed subsequently as, say, cupric ferrocyanide (Hatchett's Brown). This compound could be ignored as safely as the "iron blues," as essentially an inert solid, stable in geologic time if not exposed to sunlight. Under poorly controlled oxidation of the cyanocuprite ions (copper with four, three or two cyanides complexed to the copper), CuCN could well be formed and could not be assumed to precipitate out as a solid. The authors of the paper, by failing to include the CuCN species in their model, may have left users of the model prone to ignore the possible occurrence of the colloidal(?) form in the discharge, thus finding an effluent that failed to meet regulatory limits. (See addendum below.) This somewhat ephemeral material may well have been responsible for some large-scale failures of cyanide destruction technology to succeed at sites in North America and elsewhere. Two specific references alluding to this occurrence may be found: Griffiths et al., 1987, "The detoxification of gold-mill tailings with hydrogen peroxide," Journal of the South African Institute of Mining and Metallurgy, September, p. 282. DeVries, 1989, "Improved understanding in detoxification of cyanide leach liquors using peroxide" Rand01 Gold Conference, Sacramento CA, p. 15 1. It is advisable for workers in this area of cyanide removal or recycle to avoid assuming that CuCN will, with certainty, precipitate out of solution, or that the dicyanocuprite ion will part simultaneously with both cyanide ions and allow cupric copper to be dealt with by other reactions. One further piece of work may be incomplete. That is, a definitive study of the long-term behavior of CuCN that does precipitate when exposed to other compounds present in neutralized tailings. Addendum. The extremely complex reaction of cupric ion on cyanide- which, of course, results in oxidation of cyanide to cyanate, probably through the intermediate stage of forming cyanogen [(CN)2] from oxidation of HCN with its hydrolysis to cyanate and cyanide ion - involves two cyanide moieties and the reduction of the cupric ion back again to the cuprous state. This, then, can form the CuCN potential "solid" that was already the undesirable species. In cyanide cleanup, with either air/SO2 or hydrogen peroxide, copper is generally employed as the catalyst, so it is unclear what happens when the CU(CN)2-1 is ultimately exhausted. Certainly, the field evidence is quite clear that all

Jan 1, 2002