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By Weijian Yu, Genshui Wu, Jian Zhang, Yong Ning

Aiming at characteristics of the soft coal roadway engineering geology and deformation, the exposure research of the soft coal seam, theoretical analysis, numerical calculation and program design were carried out. First of all, according to the disclosure, soft coal seam engineering characteristics and geological conditions in Xinxing coal mine were expounded. Meanwhile, the soft coal roadway bolt supporting axial effect, the bolt supporting transverse effect, the anchorage body mechanics strengthening mechanism and problems such as supporting strength have been analyzed. The paper pointed out that the soft coal with very small cohesive force can form a composite anchorage body under the combined action of the anchor cable group, the W-type steel band and the metal net. The soft coal body in the plastic damage area formed anchorage balance arch with a certain bearing capacity and adapted to the rock deformation, and which played an effective support role in the roof outside the balance arch of the roadway. Therefore, the full-cable support technology was proposed. And then, numerical calculation and comparative analysis were conducted on three schemes of the roadway with non-support, with combined support of “anchor bolt and cable” and the full-cable support. The results show that when the roof is supported by anchor cable, it not only controls the continued increase of the subsidence value of the roof, but also effectively suppresses the convergence amount of the floor heave value and the amount on the sidewalls of the roadway. Whether it is from the stress distribution situation, or the viewpoint of plastic zone scope, the combined support of full-cable is more reasonable than that with non-support and that with the combined support of “anchor bolt and cable”. Finally, based on the features of soft coal seam in No.2265 work face of Xinxing coal mine, the relevant support schemes were put forward, and its specific support parameters and cross-section designs were drawn as well.

Jan 1, 2017

By R. Norman Holme

Sulfonate promoters and collectors were originally developed as relatively inexpensive anionic collectors to selectively concentrate the iron-bearing minerals in low grade iron ores and to remove iron and other mineral contaminants by flotation in the processing of glass sands, feldspar and other ceramic raw materials. The first reagents produced were effective collectors for such iron oxide minerals as goethite, hematite, limonite, magnetite and siderite. Further development has resulted in additional reagents with a broadened application to the selective flotation of a variety of nonsulfide and nonmetallic minerals. The products in this group are anionic sulfonate-type promoters, natural petroleum sulfonates, and synthetic modified sulfonates. Beyond their applications to iron-bearing minerals, the sulfonate collectors have gained commercial use in the flotation treatment of barite, celestite, fluorite, apatite, chromite, industrial garnet, kyanite, mica, rhodochrosite and cassiterite. in addition, extensive laboratory studies indicate their potential suitability for flotation of the following minerals: Anatase Anhydrite Bauxite Beryl Calcite Chlorite Some Copper Carbonates Corundum Dolomite Hornblende Huebnerite Ilmenite Lepidolite Magnesite Phosphate Minerals Rutile Scheelite Spodumene Sillimanite Talc Tourmaline Wolframite Zircon Sulfonate promoters are water soluble or dispersible liquids, syrups or pastes of varying viscosity with colors ranging from dark blackish-brown to clear. Customarily, they are fed to the conditioner or flotation as 5% to 30% water solutions or dispersions. In addition to their normal collector activity, certain of these products display notable frothing characteristics. Improved collector efficiency and adequate froth control can frequently be obtained by employing mixtures of sulfonate reagents, of ten with results quite markedly superior to those achieved with equivalent dosages of the individual promoter alone. Reagent requirements will vary with the type of mineral being treated; however, a general overall range of 0.13 to 2.0 kg/metric ton (0.25 to 4.0 lb/ton) of feed should cover most applications.

Jan 1, 1986

By Robert A. Jankowski, Charles A. Babbitt, Natesa I. Jayaraman

Today, approximately 31% of underground coal mining operations have designated occupations or areas with more stringent (below 2.0 mg/m3) respirable dust standards in place. Many U.S. coal mines have been able to maintain compliance with very low dust levels, and the Bureau of Mines, in an effort to identify those factors contributing to low dust levels, completed a survey of twelve continuous miner sections which were able to maintain their dust levels at or below 0.5 mg/m3. The quantity and quality of face ventilation appeared to be the most important factor in controlling dust exposure. The results directed Bureau research towards identifying improved face ventilation techniques for reducing quartz dust levels on continuous mining operations. This research has involved the evaluation of a half-curtain face ventilation technique which reduces the cross-sectional area of the entry, thus increasing the air velocity over the operator. The Bureau has also conducted extensive laboratory studies to compare tubing versus brattice, including an evaluation of both systems in a crosscut turn. In an effort to improve ventilation and reduce dust exposures of roof bolter operators working downwind of the continuous mining machine, the Bureau has evaluated two novel face ventilation systems. The first system consists of a fan discharge tubing technique which routes dusty air away from the return entry being bolted. In a technique similar to the fan/tubing concept, exhaust air from the upwind continuous miner face can be scrubbed using a dust collector to capture a portion of the dusty air, clean it, and supply it to the bolter operators as a split of clean intake air. A better understanding of face ventilation techniques should lead to solutions to the problem of more stringent dust standards due to quartz.

Jan 1, 1989

By J. Siege1, R. S. Popielak

A computer model was used to simulate the effect of leakage through confining beds on a hypothetical in situ leach operation in an aquifer with a thin mineralized section. The simulations were used to assess economic and potential environmental implications of leakage on an in situ leach operation. Four scenarios were modeled - three representing cases with different degrees of confinement, and the fourth representing a stratified ore zone under a low degree of confinement. The model simulated the travel path and travel time of lixiviant given the hydraulic conditions prescribed in each scenario. The results show that the travel path and travel times vary by about 10% for confinement ranging from nearly impermeable (essentially no leakage) to an extremely low level (thin confinement with a permeability about equal to that of the aquifer) for isotropic aquifer conditions. The aquifer thickness contacted by the lixiviant varied from 85% to 90%. Horizontally, 55% to 60% of the aquifer was contacted. The flow paths in the stratified aquifer were reduced in the vertical direction and expanded horizontally. Forty- five percent of the aquifer thickness was contacted vertically with 88% contacted horizontally. Break- through time for lixiviant traveling at mid-depth of the aquifer increased two times when the leakage was increased from nearly zero to a value representative of thin and highly permeable confinement. Leakage affects the economics of in situ mining by (1) increasing the time required to leach out the ore, and (2) increasing the thickness of the ore zone aquifer to be restored. The potential for environmental impacts appears to be minor. In all four simulations, the lixiviant did not contact the confining strata. Restoration would be limited to cleanup of the ore zone aquifer within the thickness contacted by lixiviant and within the radius of the lixiviant that traveled farthest away from the production cell.

Jan 1, 1988

By Brian Flintoff

SCOTT COULTER: Our second speaker today is Brian Flintoff. Brian graduated from the University of Alberta in 1973 with a master's of science degree and worked as a project metallurgist at Brenda Mines until 1977. He returned to the University of Alberta in late 1977 to work on his Ph.D. and later joined the faculty of engineering in 1980, leaving in 1988 to rejoin Brenda in the newly formed process technology division. In 1998 he left Svedala-Brenda to take a position with Simons International's newly formed enterprise development services group. His topic today is "People: Reduce, Reuse, and Recycle." INTRODUCTION BRIAN FLINTOFF: Historically, the mining industry has relied on the people within a corporation to get things done. The culture of most companies was one of a slow but sure development of people, processes, equipment, and technology. Many of the larger companies supported in-house engineering and research organizations. However, the metal market pressures of the 1980s and the emergence of the global economy in the late '80s and early '90s put enormous financial pressures on the industry. As the old adage says, "necessity is the mother of invention," and the mining industry, like many others, has undergone a transformation. A cornerstone element of the transformation was the downsizing (or right-sizing) that was calculated to "trim the fat" from the capital and operating budgets. The goal was and is to find the highest sustainable operating profit.' While I'm sure one can find examples of losers (e.g., those who can be classified as suffering from corporate anorexia or institutional Alzheimer's) in such business reengineering efforts, there is little doubt that, on the whole, the transformation has been successful. The reality that now faces mill management is "How to do more with less?" At the risk of oversimplifying, one can look at this on two levels:

Jan 1, 1998

By L. Adler

INTRODUCTION The purpose of this paper is to describe in engineering terms the singular and basic operational kinematics of all excavating and bulk handling equipment. The safe and efficient use of equipment will be fostered by enabling the engineer to conceptually duplicate the operators' "tasks" in directing and controlling it. A simple graphical and notational scheme will be introduced to define the essential gross repetitive motions. The significance of this effort includes filling a vital analytical gap at the work face, and in establishing direct communications between the engineer and equipment operator. Since all types and sizes of earth- moving equipment are covered, a major concern has been to integrate equipment specialties into d single encompassing discipline. Equipment types are closely related to the working conditions, and their selection and evaluation are also considered. The extension to human factors analysis is natural and predictable. Since human [actors are involved, the operational analysis can also be considered a catch-all for many subtle vagaries. BACKGROUND Symbolic modelling linked to a notational system will be developed and intensively used. This constitutes the "language of science," providing a shorthand method which gives immediate recognition, is fluent, explicit, reinforcing, unequivocal and provides limitless associations. Many examples of symbolic modelling exist in chemistry, physics, the military, engineering sciences and it is a major tool in mapping. Its use is so ubiquitous that it often goes unnoticed. On the other hand it must be realized that something is lost in simplifying and idealizing what is an extraordinarily complex occurrence. The components describing the basic operational kinematics of equipment include: (1) the vehicular motions as a rigid body and (2) the job-end motions focussing on excavating (Adler, 1985). These provide all of the components in the load-haul-dump sequence (L-H-D) , and will yield three dimensional expressions. Complex patterns can also be assembled from these simple elements. An existing equipment designation employed in earthmoving will be integrated into this approach (Carson, 1961). Along with the basic motions, corresponding working conditions will be noted.

Jan 1, 1986

By William S. Eaton

Managing a mine or mineral processing operation in today's environment presents many unique challenges to mining companies. In recent years, headlines around the world have publicized major incidents such as those associated with the Marcopper Mine in the Philippines, the cyanide release in Guyana from the Omai gold mine and numerous other mishaps associated with mining. These notable upsets were due to engineered system failures, either during normal operations or as a result of uncontrollable events. There fore, many mining firms are attempting to assess their overall risks and establish some priority for expenditure of limited operational and environmental budgets. Failure of an engineered component can have dire operational consequences that affect profitability until repair or replacement can be completed. In addition, there may be significant environmental consequences associated with the engineered system failure. This, too, may affect profitability for an extended time or even the viability of an operation. An engineered risk assessment (ERA) can support a mining company's efforts to cost effectively manage environmental risks associated with the operation of its facilities. The ERA approach also provides information about the current condition of engineered systems. This ERA approach provides a systematic review of the risks of an engineered system failure and the possible associated environmental consequences at mining company facilities. This approach has been successfully applied to many industries and types of engineered systems (Barneich et al, 1996; Whitman, 1984; Power and Tomkins, 1974; Keeney, Nair, and Kulkarni, 1979). This approach evaluates all mining and mineral processing facilities located at a single site and compares them with similar facilities located at multiple mine sites. The ERA approach has been applied to a mining, milling, smelting, leaching and refining operations. These assessments can be part of a comprehensive environmental program to aid in the recognition and cost effective management of the potential environmental risks associated with the operation of facilities. ERA results support plan-

Jan 1, 1997

By Bruce K. Cantrell

As part of a continuing study of diesel exhaust aerosol characteristics, the Bureau of Mines has applied Chemical Mass Balance (CMB) model source apportionment analysis to evaluate a size selective sampling method for determining the separate diesel exhaust and coal dust aerosol concentrations in a diesel equipped coal mine. Size selective sampling was used under the premise that diesel exhaust aerosol dominates the submicrometer aerosol size range, while most of the mineral dust aerosol is greater than 1 m. A micro-orifice, uniform deposit impactor (MOUDI) was used as the size selective sampling method. Concurrent samples of mine aerosol were collected using a dichotomous, respirable aerosol sampler. This sampler separated the fine and coarse size portions of its sample at 0.7 m. Trace element concentrations for in-mine aerosol sources and the dichotomous samples used by the CMB model were measured using Instrumental Neutron Activation Analysis. The CMB analysis, applied to both the fine and coarse fractions of the respirable mine aerosol, yielded the mass concentration contributed to each by diesel exhaust and mineral dust aerosol sources. The MOUDI size distribution data were analyzed using nonlinear regression to parameterize the fine and coarse fractions of the sampled aerosol and to estimate the contributions to each fraction from both coal dust and diesel aero- sol sources. For the diesel mine visited, analysis of the MOUDI samples yielded a value for the fraction of the fine aerosol attributable to coal dust of 15[]7% fine fraction coal from analysis of the MOUDI samples versus 25[]4% from the CMB analysis. The remainder of the fine aerosol in the mine was diesel exhaust aerosol. CMB analysis confirms the original premise for such sampling. After correcting for the presence of coal dust in the fine aerosol fraction, analysis of the MOUDI size distributions yielded estimates of respirable mineral and diesel exhaust aerosol concentrations that were accurate to with-in 10% of values determined by the CMB apportionment.

Jan 1, 1987

By Pieter Barendsen

INTRODUCTION One of the main problems in tunneling without the use of explosives lies in the development of tools which, at an economically acceptable level, are capable of continuous breakage of the rock, resulting in a fragmentation suitable for a smooth, uninterrupted transportation of the muck away from the tunnel face. Tools employed on tunneling machines have, so far, always been of the mechanical kind which break up the rock by a crushing and shearing action. Other methods of rock breaking are possible but, even if some of them have reached the laboratory testing stage, nearly all of them are still too "exotic" to be of any practical use in the immediate future. One must admit, however, that some highly interesting results are beginning to be reported from the USA on the use of high pressure water jets for the destruction of rock through erosion, a field pioneered in Europe(USSR and Great Britain). Mechanical breakage of the rock is effected by inducing stresses exceeding its compressive/shear strength, by loading it with a wedge of cone shaped tool, until cracks are formed and chips loosened. In order to allow the tools to act upon the rock face continuously, they are commonly shaped as rotating bodies, "roller bits" or "disc cutters", spinning freely and mounted on a revolving boring head of tunnel diameter. The thrust of the tools against the face is generally exerted in a direction which is parallel to the tunnel axis.

Jan 1, 1997

By T. Ertekin, R. V. Ramani, W. Sung

The emission of methane in the coal mine working faces has two immediate impacts with regard to coal mine planning. Firstly, at the planning stage, the coal mine ventilation system must be designed such that airflows in the faces keep the concentrations of the gas in faces areas to below one percent. If the concentrations were greater than one percent, the production machine has to be shut down resulting in a loss of productive time. In an extreme case, concentration can reach explosibility limits which may lead to ignitions and explosions. Secondly, in excessively gassy coal seams, several alternatives can be applied to the mine layout configuration, and the mining method to be predicted for the extraction of coal, so as to insure a productive and safe operation. In this study a two-phase, two-dimensional finite-difference model was used to predict the methane emission rates into active mine working area under several methane drainage schemes. The numerical model utilizes a recently developed multi-mechanistic flow theory which describes the flow of methane in an heterogeneous and anisotropic coal seam. Transport of methane in the micropore structure of the coal is described by the Fick's law of dif- fusion under the influence of existing concentration gradient, while flow in the more permeable butt and face cleats (macropore spaces) is formulated by Dar- cy's law under the influence of existing pressure gradient. The desorption of the methane is accounted with the aid of a quasi-steady kinetic sorption model. A series of simulation studies were conducted in a systematic manner by sifting through the parameters associated with the coal seam characteristics and the degasification strategies in order to understand those which most significantly affect the methane emission rates in the face areas during mining. Along these lines, the effects of coal seam properties such as thickness, porosity, permeability, sorption characteristics and scheduling of drainage process against the mining operation have been isolated to understand the role of each individual parameter on the methane emission rates.

Jan 1, 1987

By Kaye P. Strong, Desmond M. Levins, Anthony G. Fane

1. INTRODUCTION The safe disposal of tailings is the most important environmental question facing the uranium mining industry today. This is because tailings present a major long term radiological hazard as they contain about 71% of the radioactivity originally in the ore. The magnitude of the problem can be appreciated by considering the quantity of tailings involved. For example, projections of milling programs in Australia indicate that there will be 100 million tonnes of tailings produced by the year 2000 (Levins 1980). The United States currently has in excess of 100 million tonnes and by the year 2000 will have between 900 and 1700 million tonnes (Landa 1980). Present options for tailings rehabilitation include leaving them above ground, returning them to the mine or placing them in specially excavated pits. In any event, the tailings will be covered with earth and revegetation attempted. This will remove the problems of airborne radioactive dust and water erosion of the tailings, at least in the short term, but diffusion of radon-222* from the tailings may still occur. The Atomic Energy Control Board of Canada (1981) has proposed a maximum allowable radon flux from rehabilitated tailings of 0.074-0.37 Bq m -2s-1(2-10 pCi M-2 s-1); the exact level is to be determined "after discussion with other regulatory agencies and the industry". The United States Nuclear Regulatory Commission (USNRC 1980) has been more restrictive in recommending a maximum flux of 0.074 Bq m -2s-1 (2 pCi m-2s-1). As a guideline they also suggested a minimum depth of covering material of 3 m to achieve this surface flux for tailings typically found in the western milling regions of the United States. The design of rehabilitation schemes is aided by a knowledge of radon transport through tailings piles. In this paper we present the results of a theoretical analysis and experimental study of radon diffusion through selected Australian tailings and cover material. The minimum cover criteria of 3 m is reexamined in the light of these findings. [ ]

Jan 1, 1981

By R. L. Grayson, G. Mishra

Introduction Many old underground mines have been developed extensively into reserves requiring large numbers of workers to maintain and rehabilitate lengthy track haulages and/or belt conveyor systems. Often, large numbers of fans are required to maintain minimal airflows in areas where poor conditions and huge leakages prevail. Problems that beset one such coal mine and projects undertaken to restore it to potential profitability are presented. Brief history The Vesta No. 5 mine was opened in 1908. Drift development began near the Monongahela River, which was the primary avenue for shipping metallurgical coal to coke ovens in Pittsburgh. Over the ensuing 75 years, the main haulage entries were developed a distance of 16.1 km (10 miles) to the present mining area. Track haulage was used to transport coal from working sections to the river loading facility. As haulage distance and areas of high water influx increased, management decided, in 1973, to construct an underground conveyor belt system. The project spanned 36 months and culminated in a system of 22 flights, ranging from 205.7 to 1767.8 m (675 to 5800 ft) long. Six intake and eight return entries were developed as the mains were advanced. These entries provided adequate ventilating capacity over the years, but as extensive gob areas progressed and roof conditions deteriorated, the entries became inaccessible and more resistant to airflow. By 1981, five fans were being used to circulate 943.9 m3/s (555.5 cfm) of air through the mine. An average of 6440 t (7100 st) were produced from 20.8 continuous miner unit shifts and 2.7 longwall unit shifts each day in 1981. The ratio of support to production personnel was 1.94. To maintain the outby 16.1 km (10 miles) of haulage, 110 personnel were used. The average cost per ton was $43.54 ($39.58 per st), resulting in excessive bottom-line losses. The mine possessed a rather promising reserve, mostly in lease, of good quality, medium-sulfur metallurgical coal, but it also had a number of formidable operating problems. Operating problems Contributing factors toward relatively low productivity and high costs were the conditions of the track haulage, conveyor, and ventilation systems. Productivity was affected greatly by downtime associated with the conveyor system, which lacked adequate access for making splices or other repairs. Many areas within the conveyor belt entry itself required more

Jan 1, 1988

By Phoebe L. Hauff, Jeffrey P. Kurtz, Paul D. Chamberlin

Clay minerals can be a source of problems in heap leaching operations. The ore for many heap leach operations is commonly acid-sulfate alteration zone rocks which may be found in hot springs deposits as well as the supergene zone of other near- surface gold deposits. Clays can be important constituents of these ores. Swelling clays, such as smectite, can come loose from the agglomerated ore, migrate within the heap, concentrate in pore spaces and often plug up flow channels in the heap Cement is often added to the ore as a binder during the agglomeration process. Lime is added for pH control. Their presence, combined with the sodium cyanide commonly used to leach, can raise the pH of the leach solution to 10 or even 12. This very alkaline solution can cause dissolution of some mineral Some clays, such as kaolinite, are not very stable in high pH solutions. They can be dissolved and reprecipitated as other clay species if sufficient silica and the appropriate cations are available in solution. Silica gels and amorphous materials can also be created from the dissolved kaolinite and other minerals and act as retardants to the flow of leach solutions in the heap. Therefore, before a heap leach operation is built, the operator should evaluate the clays in the ore and test the agglomeration procedure. It is important to know not only the mineralogy and the chemistry of the leach environment but how the minerals and solutions interact and what the products of this interaction are. A hypothesis is presented in this paper that expanding smectite- bearing clays can be generated within the heap. How and why this happens will be explained.

Jan 1, 1987

By Bernhard C. Koch, Richard W. Hutchinson, Bernhard Free

The Tertiary Silverton caldera collapsed asymmetrically to the southwest 27.5 m.y. ago. Resurgent volcanic activity is expressed by a quartz-monzonite stock at the southern caldera margin and by the Red Mountain porphyry stock in the northwest of the caldera which were emplaced at 25 m.y. and 11 m.y. b.p., respectively. The apical Eureka graben formed between 27.5-22.5 m.y. ago and transects the Silverton caldera southwestward. In the northeast it terminates at the 22.5 m.y. Lake City caldera margin. This graben as well as caldera faults provide dominant structural control for base and precious metal deposits. The epithermal gold mineralization of the Gold King mine is proximal to the Red Mountain stock in the north central part of the caldera. Here, west-northwest striking faults related to the collapse of the caldera intersect northeast faults of the Eureka graben. Superimposed upon earlier base metal mineralization at 23 to 22.5 m.y. b.p., the 11 m.y. younger gold metallization was a distinct event at Gold King. Supergene enrichment of the hypogene precious metal assemblage resulted in a metal tenor of up to 3,016 g/t gold and 1,088 g/t silver. The deposit is hosted by quartz-latite tuffs, flows, and flow breccias of the Burns Formation. Veins, up to 5 m wide follow steeply southeast and northwest dipping structures of the Eureka graben. Subsidiary veins up to 2 m wide branch off the main lode and are termed "flat veins" because of their conformable strike but shallow dip of 45' to 50' toward the northwest. Meteoric water-dominated hydrothermal solutions generated by the Red Mountain stock ascended through intensely fractured, ferruginous

Jan 1, 1988

By G. H. Lewis, C. L. Krcil, N. S. Harding, D. E. Havekotte

INTRODUCTION Limestone injection work was completed in a 440-kW (1.5-million Btu per hour) pilot-scale combustor in preparation for a 50,000-kg/h (110,000-lb per hour) industrial steam boiler limestone injection test program. The main objectives were to: (1) identify the important variables that affect sulfur capture, (2) determine the maximum SO2 removal, and (3) determine the effect of limestone addition on furnace operability and heat transfer. A more detailed report is given by Lewis et al. (1985). The test program was divided into two phases. The objectives of Phase I were to determine the important variables that affect SO2 reduction and to investigate the effect of limestone addition on combustor operation and heat transfer surfaces. The objective of Phase II was to determine, using only the important variables identified during Phase I, the conditions that resulted in the greatest sulfur dioxide reduction. Phase I Test program Using one limestone, five variables were identified as having an effect on SO2 reduction. These are noted with their respective ranges: heat release rate [103 to 166 kW/m3 (10,000 to 16,000 Btu per hour-cu ft)], Ca/S molar ratio (0 to 3), flue gas oxygen concentration (3% to 5%), additive particle size [90% -150 µm (-100 mesh) and 90% -45 µm (-325 mesh)], and burner stoichiometry (SR = 0.75 and >1.15). A 2.3% sulfur northern West Virginia coal, pulverized to a nominal grind of 70% -75 µm (-200 mesh), was burned for the tests. The limestone used was obtained from a Maysville, KY, quarry and contained 51.75% (wt) CaO. An 11-run statistically designed program (plus three duplicates) was used to screen the five variables. The effect of limestone injection on the combustor heat transfer surfaces was determined by comparing the heat transfer rates through a slag panel and fouling probes during periods with and without limestone injection. Combustor operation was determined by noting any operating problems that occurred during the tests. Test procedure After furnace wall temperatures had stabilized at operating conditions, heat transfer surfaces (slag panel and fouling probes) and the convection section were blown clean, and a four-hour test was begun. The

Jan 1, 1987

By R. Fedel-Moen, R. W. L. Kimber, B. D. Williams, S. R. Ragusa

The use o f cyanide to extract gold from ores is widespread. In the process of extracting gold, cyanide may form metal cyanide complexes with a variety of other metals present in the ore. This study was carried out to investigate the biodegradation of the Cu(I)-cyanide complex and tetracyanonickelate under varying pH and nutrient additions. Detailed studies using shake flask cultures and bioreactors showed that naturally occurring bacteria residing on gold-leach-pad ore uti¬lized the Cu(I)-cyanide complex (CuCN) and tetracyano-nickelate, Na2(NiCN4), as carbon and/or nitrogen sources. Ammonium-nitrogen (NH4+-N) and/or cyanate (CNO-) accumulated in solution as the metal cyanide complexes were degraded. The accumulation of cyanate as a product is a concern due to its toxicity. Degradation rates were increased if a growth substrate (peptone, 0.1%) was added to the medium or if bioreactors were primed with peptone before addition of the metal-cyanide complex. Tetracyanonickelate was degraded at pH 8 without added peptone, but it was not degraded at pH 10. The addition of peptone was required for degradation of the Cu(I)-cyanide complex at pH 8 and pH 10. At pH 8, the addition of peptone resulted in the accumulation of cyanate, whereas, at pH 10, cyanate was not detected in the cultures. Biodegradation of metal-cyanide complexes was faster (approximately 1.5 to ten times faster) in bioreactors than in shake flask cultures. The addition of peptone was not required for degradation of metal-cyanide complexes in bioreactors. However, priming the bioreaetor with peptone resulted in higher metal-cyanide degradation rates over multiple cycles.

Jan 1, 2000

By Patrick J. Ryan, Johann F. Kerl, Angelos N. Findikakis

Introduction Tailings disposal methods used by the mining industry include the use of on-land impoundments, surface and underground mine backfill and marine disposal. The latter method, where applicable and under certain circumstances, can be the most economical and best environmental option. The term marine disposal is used to describe deposition of tailings in the open ocean, coastal waters, fjords and inland lakes. A major element in planning and designing a marine outfall is the assessment of its potential environmental impacts. The nature of the disposed tailings, and the physical, chemical, and biological characteristics of the receiving water body must be carefully assessed prior to deciding whether marine disposal is an acceptable option for a particular project. During the last decade, the North American mining industry has demonstrated a commitment to studying potential adverse environmental impacts of marine tailings disposal and to continue monitoring of the receiving waters. Extensive monitoring programs have been undertaken by Canadian mines. Marine tailings disposal is not used in the United States today. One of the largest U.S. mines proposing to use this option is the Quartz Hill mine of U.S. Borax in Southeastern Alaska. To assure the regulatory agencies and the public of the viability and acceptability of this option, U.S. Borax undertook an extensive program of environmental studies aimed at assessing the impact of marine tailings disposal. This program was part of the feasibility and environmental permitting studies conducted by Bechtel Civil & Minerals, Inc, on behalf of U.S. Borax. Due to the unique nature of the proposed tailings disposal system, a Board of Consultants was established. The Board assisted in planning the modeling and field program and reviewed ongoing work. This paper outlines the basic elements of the marine tailings disposal study program which can be viewed as a prototype for similar studies elsewhere. Background The Quartz Hill molybdenum deposit is located on a knoll in an elevated valley in a mountainous region of southeast Alaska, about 45 miles east of Ketchikan (Figure 1).

Jan 1, 1986

By M. P. Hardy, C. J. H. Brest van Kempen, H. N. Maleki

To evaluate the influence of uniformly tensioned roof bolts on roof stability, mechanically anchored roof bolts were installed using the conventional bolter and a specially developed thrust and torque controlled retrofitable box mounted on a bolting machine. Thirteen thousand bolts were installed over 19 crosscuts in a 4 heading development in an operating coal mine. Geological , bolt and mine parameters were recorded at bolt installation, and bolt tension histories, roof sag and other changes evident in the rooms, crosscuts and intersections were recorded over a 2 month period. In all, 24 variables and 11 responses were quantified at 67 bolt clusters. These data were analyzed statistically to determine the significant variables influencing roof stability. As the mine development was safe and stable during the monitoring period, roof sag after two months was used as an index to quantify "stability." The variables of significance were shown to be coefficient of variation of bolt tension, bolt anchorage capacity, roof lithology, roof shape, standard deviation of bolt tension, and dip of roof lithologies. It was concluded that of the variables with which the mine operator has control, bolt tension uniformity was the most significant parameter at this mine. Other variables, such as intersection type and time between mining and bolting , were not significant, whereas, some other potentially significant variables such as horizontal stress or roof joint frequency did not vary sufficiently for their influence on stability to be evaluated.

Jan 1, 1986

By P. Somasundaran, F. F. Aplan

The Formative Years Douglas Winston Fuerstenau was born December 6, 1928 in Hazel, SD to Edwin Arnold and Hazel Pauline (Karterud) Fuerstenau. His father, of German extraction, was a farmer and, later, a cabinet maker, while his mother, of Norwegian lineage, was a school teacher. His parents are still alive today at ages 91 and 88, respectively. Two younger brothers, Richard and Maurice and a sister, Jean, completed the family. The family first lived near Watertown and later in Lemmon and Mobridge, SD. The 1930's were a difficult period for any family in South Dakota. A decade-long drought made this a part of the dust bowl, which together with a serious infestation of grasshoppers (the locust plague described in The Bible), and the world-wide depression, made living difficult for all. In spite of these privations, Doug has many happy memories of his childhood which can be attributed to the resilience of youth and a strong family structure. He has fond memories of summers spent with his cousins on family farms where the children were introduced to farm chores at an early age. When he was old enough, he worked in a grocery store after school. At the age of 14 he spent the summer hearding 2,000 sheep near Lemmon, SD as his contribution to the war effort in that labor-short period. He remembers well the dust storms of the 1930ts, the destruction at his grandmother's farm by tornado, pulling weeds on the farm, threshing grain with a monstrous steam-driven machine and such pleasant childhood activi- ties as fishing for bullheads, hunting for rabbits, searching for arrowheads in newly plowed fields, and ice skating in the winter. Early in life his parents instilled in him a strong work ethic, a love of learning and the need for a good education. In 1943, the family moved to Rapid City, SD. With a population of 13,000, it was

Jan 1, 1989

By Charles U. Keranen

The Tilden Mine is an iron ore mining, processing and pelletizing facility located on the Marquette Iron Range in Michigan's Upper Peninsula. It has a designed capacity to produce 8 million long tons of pellets annually from 21 million long tons of crude. It is a partnership owned by The Cleveland-Cliffs Iron Company, Algoma Steel Corporation, LTV Steel Corporation, the Steel Company of Canada, Wheeling Pitts- burgh Steel Corporation and the Sharon Steel Corporation. Cleveland- Cliffs manages and operates the Tilden. The crude ore averages 35% iron and is primarily non-magnetic iron oxide and silica. The major ore mineral is martite. Hematite, earthy hematite, goethite and magnetite are also present. The principal gangue mineral is chert. The ore grains are mostly in the 10 to 25 micron size range which necessitates very fine grinding for liberation. The concentrating process is the most complex one used in the iron ore industry. The unique technology for Tilden was developed in the mid and late 60's. The design and construction occurred in the early 70's and operation of the first phase which was of 4 MLTPY capacity began in October. 1974. Expansion of the facility to the current 8 MLTPY capacity was completed in 1979. Chemicals are the key to the beneficiation process at the Tilden Mine. The flowsheet utilizes selective flocculation - desliming, cationic silica flotation and concentrate dewatering by thickening and filtering. Nine different process chemcals are used. The control of the reagent feed systems is highly auto- mated. The entire concentrator is operated from a central control room. The concentrator flow diagram is presented in figure 1. The plant consists of 12 primary autogenous grinding lines, a common crushing circuit to handle an intermediate product within the primary grinding section, 24 pebble mill lines with cyclones and deslime thickeners, 12 flotation lines, 4 concentrate thickeners and slurry tanks, 8 lines of filters with six machines per line, and 2 tailings thickeners. A pumping system was installed in 1980 to permit the direct return of tailings thickener overflow water back to the reuse water pumphouse.

Jan 1, 1986