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By A. S. Winters

INTRODUCTION Homestake Mining Co. owns and operates the Bull¬dog Mountain mine near Creede, CO. Lead-silver con¬centrates are produced from ores extracted from narrow veins which dip from 1.04 to 1.39 rad (60 to 80°). An¬nual mill throughput is approximately 90 700 t (100,000 st). Ore shoots average 2 m (7 ft) in width and the ore is often weak, muddy, and poorly consolidated. Walls can be either strong or highly fractured along the strike of the veins. Cut-and-fill methods were initially selected to mine the deposit and considerable overhand mining occurred during the initial stages of the mine. Because of un¬stable ground conditions, mine management recognized that other methods would have to be employed. Undercut-and-fill mining was adapted to extract the loose ore from narrow veins and is presently the primary method used. Approximately 80% of the annual pro¬duction is extracted by this system. Loading ramps are positioned along the strike of the vein on 90-m (300-ft) intervals to allow for 45-m (150-ft) long cuts in each direction from the extraction raise. This interval was selected for efficient slusher operation and stope cycling. Loading ramps rather than chutes were selected for ore movement into 4.5-t (5-st) Granby-type cars because of extremely sticky ore conditions. Mined-out areas are filled with deslimed mill sands. Sand is pumped from the mill into two mine storage dams where the water is decanted. Each dam has a 544-t (600-st) capacity. Sand is pumped into the mine at a 30 to 40% slurry through 2195 m (7200 ft) of 51-mm (2-in.) thick wall pipe. When called for, sand is reslurried within the dams to a density of 55 to 65% and pumped to the particular mine opening requiring fill. Dry cement is added to the slurry prior to pumping. The capacity of the mine fill system is 20 t/h (22 stph). DEVELOPMENT Vertical development below the main adit level oc¬curs on 61-m (200-ft) intervals from an internal winze. Main haulage drifts or laterals are 2.7 x 2.7 m (9 x 9 ft) in section and driven parallel to the vein structures. Ground conditions generally dictate that better progress can be realized in the footwall than in the hanging wall. A distance of approximately 21 m (70 ft) is maintained between vein and drift openings to accommodate a smooth 18-m (60-ft) radius curve for crosscuts which are driven on 91-m (300-ft) intervals. Approximately 18 m (60 ft) of tail room is driven past the vein inter¬section to accommodate train loading. Ore and waste are hauled using diesel or electric driven locomotives operating over 18-kg (40-Ib) rail [762-mm (30-in.) gage]. Loading ramps are established above each crosscut¬ vein intersection. Scram drifts are then driven in the vein to determine the width and grade of the ore. If a prior decision to mine the block had been reached, then the scram is driven only far enough [about 12 m (40 ft)] to establish a drawpoint and scraper tail room. In most cases, the drawpoint is a 2.1 x 2.1-m (7 x 7-ft) cutout driven about 2.4 m (8 ft) into the footwall. On the level above, a raise bore station is established in the footwall to one side of the crosscut-vein inter¬section. The station is of sufficient size to accommo¬date a Robbins 41R raise boring machine. Support for the station is accomplished with 1.2 and 1.8-m (4 and 6-ft) roof bolts and wire screening. Once completed, a 152-mm (6-in.) concrete slab is poured and the raise drill set in position to center the 1.5-m (5-ft) diam hole approximately 1.8 m (6 ft) from the vein footwall. A 229-mm (9-in.) pilot hole is then drilled to the scram cutout below and reamed back to full diameter (see Fig. 1). STOPING Initial mining in a typical undercut-and-fill stope be¬gins from the top of the borehole. The initial cut can be either immediately beneath the track level or the old scram drift depending upon the particular situation. Cuts vary but are generally 2.4 to 3 m (8 to 10 ft) high. Support is accomplished by square-set timber, overhead caps, or roof bolts and screening depending upon ground conditions. Once the initial cut is removed, underhand timber and screen is installed and cemented sand fill placed. The stope is then ready for undercut mining. Prepara¬tion work begins with sinking 4.5 to 6 m (15 to 20 ft) on the borehole plus cutting a slot to the vein. As sinking progresses, hanging sets are installed to provide man¬way and service facilities. Hanging sets 1.7 m (5 ft 8 in.) long and 1.6 x 2.2 m (5 ft 4 in. x 7 ft 4 in.) in section are installed with steel hangers similar to shaft sets. Framed 203 x 203-mm (8 x 8-in.) timbers are made up in advance and shipped to the stope during sinking operations. When manway and service facilities are completed to the floor of cut to be mined, generally 3.65 to 4.3 m (12 to 14 ft) below the previous sand fill, the ore is removed to the hanging wall plus a round each way along the vein. Levelers or long stringers are then installed over the raise opening to support the work deck, slusher, and grizzly opening. All waste and ore from the sinking process is fed by gravity to the raise (see Fig. 2). Once the stope is equipped, mining pro¬gresses outward in each direction from the grizzly open¬ing. Stope equipment consists of 907 kg (2000 lb) pull air tuggers, timber skips, 11-kW (15-hp) ventilation fans, 15 to 22-kW (20 to 30-hp) electric slushers, 762 to 1270-mm (30 to 50-in.) scrapers, 67-mm (25/a-in. ) jackleg drills, as well as other drilling and blasting supplies. In narrow stopes [generally less than 2 m (7 ft) wide], no support is required for the underhand caps of the previous fill. Where wide sections in the vein occur, segment braces are installed beneath the caps to insure that the sand support cannot slip away from the hanging wall. Sections as great as 6 m (20 ft) wide and 4.5 m

Jan 1, 1982

By P. Corser

Operational and geotechnical constraints on mining within steeply dipping deposits and historic landslide areas, which are characteristic of subarctic coal deposits in Alaska, will be discussed. The current plan for mining through these areas will be presented, as well as an alternative plan for minimizing the adverse effects of mining through an historic landslide area.

Jan 1, 1988

By J. Chia, M. H. Kobler, J. Cheung, M. J. Robison

INTRODUCTION The Lake Merced Transport is being constructed as part of San Francisco's Clean Water Program. The Clean Water Program, which began in the early 1970's, is designed to control wet weather, storm sewage overflows into the Pacific Ocean and San Francisco Bay. To accomplish this goal, a system of large underground storage and transport facilities, pump stations, modernized treatment plants, and outfalls is being built. By 1991, eighty percent of the system had been completed and was in operation. One of the remaining projects for the Clean Water Program is the Lake Merced Transport. The Lake Merced Transport is located in the extreme southwestern portion of San Francisco, as shown in Figure 1, Project Location. The transport will convey dry weather flow from the existing Lake Merced Pump Station to the Westside Pump Station. During wet weather, the transport will fill with sewage and runoff, with a storage capacity of 36 million liters. The Lake Merced Transport was expected to be excavated entirely within the Colma Formation, an alluvial deposit consisting of weakly cemented sand. Once excavation of the access portal at the Sloat/Great Highway parking lot began, it became evident that a portion of the tunnel would be excavated through beach/dune sand. Based upon the expectations of the Geotechnical Design Summary Report (GDSR), encountering beach/

Jan 1, 1993

By J. F. Kvitfkovich, R. L. Bate

Coal production in the Central Appalachian region has a long history, and it remains a major US coal-producing region. The coal industry consensus is that the region is in a state of decline as evidenced by the recent lower regional output. Are Central Appalachian coal reserves adequate for the region to continue as a major supplier of low sulfur coal? To answer that question, this paper focuses on the Central Appalachian coal reserve base estimates and production trends. Central Appalachian (CAPP) reserve evaluations face the difficulty that there are no publicly available databases that provide a thorough analysis of remaining CAPP reserves. Based on best available Energy Information Administration (EIA) data, a coal reserve estimate of 17.6 billion tons is developed. This estimate appears high according to industry perception. Recent United States Geological Survey (USGS), and other state agency coal reserve estimation projects are evaluated as alternative approaches to the re-serve estimation problem. Using the conclusions from these studies, the CAPP reserve base is estimated at 7.1 billion tons. Another perspective on the CAPP reserve base is provided by reserve holdings of major CAPP operating companies reported at 5.2 billion tons. A conceptual mining cost analysis indicates substantial price increases will be necessary to offset this deterioration in the economic viability of the reserve base.

Jan 1, 2004

By K. Holmberg, D. Herriot

This paper by Mr. McIvor is yet another contribution to the eternal controversy surrounding the benefits, or otherwise, of high circulating loads in closed circuit grinding. Throughout the paper it is difficult to establish whether the author is referring to effects relating to the mill, classifier, or to the overall circuit. Table 1 is put forward as indicating the advantage of operating at high circulating loads. From the two (it would be interesting to see more) product sizing points provided in each case, the conclusions to be drawn are that cases one and two have very similar size analyses. The third case, however, indicates a much flatter curve, that is to say, higher fines production and a very wide spread of size ranges. If the objective of this circuit is to provide a closely sized product with minimum overground slimes to the downstream process, then it would appear to be the less "efficient" by far. No data is given for product pulp densities either. The term "efficiency" should never be used without a careful and clear definition. In this case, "efficiency" is used to express an apparent percentage reduction in operating work index (Bond) from the measured "standard" at 250% circulating load operation. The Bond calculation is itself based upon

Jan 1, 1990

By Jean-Michel Rendu

The semivariogram is one of the most important tools available to the geostatistician for the analysis of mineral deposits. To be complete, a semivariogram study must include a minimum of three parts which are highly interdependent. These are: calculation of the semivariogram, mathematical modeling, and geologic interpretation. The known factors and conditions which controlled or influenced the distribution of mineralization within the body being studied should be considered in the analysis .of the semivariogram. The presence and characteristics of successive phases of mineralization, the chemical and physical properties of the host rocks and the mineralization should all be considered. This requires close and continued communication between the geologist and the geostatistician prior to and during the semi-variogram study. Geologic stationarity is needed if meaningful semivariogram results are to be obtained. The first step is, therefore, to determine whether the mineralization can be analyzed as a whole or if different geologically homogeneous sections must be defined and evaluated separately. Different methods can be used to calculate an experimental semivariogram from known sample values. The method chosen will depend not only on the anticipated and actual statistical properties of the deposit, but also on its geologic characteristics.

Jan 1, 1982

By Timothy F. Markey, Brian C. Strohm

This paper presents a case study which applies risk assessment techniques in evaluating potential environmental problems associated with past waste disposal practices at a 65 acre non-ferrous metals and metal alloys manufacturing facility. Waste management activities at the site involved a number of surface impoundments for gangue material, neutralized acid rinse waters, acid rinse sludges, demolition debris, storm water runoff, stockpiled ores, and a permitted residual waste landfill. As a result of these activities, the She contained 9 identifiable waste management areas impacted by over 35 potentially hazardous chemicals and metals. Extensive environmental data obtained from investigations at the Site revealed heavy metal and volatile organic chemical contamination in various environmental media. These data were utilized to evaluate potential risks to human health and the environment. The human health risk assessment estimated potential health risks to workers, residents, and Site trespassers under current and expected future uses of the Site. Areas of the Site were identified which were considered to present potential health concerns under the various exposure scenarios. Results from the human health risk assessment were utilized to develop cost effective risk based soil remediation goals protective of the public health and environment.

Jan 1, 1992

By R. J. Sweigard, R. V. Ramani

The effects of postmining land use planning are addressed in this paper. These effects may include savings in reclamation costs and increased value of reclaimed land. However, this may result in greater planning costs and possible delays in production. Therefore, an economic analysis was undertaken in an effort to determine the importance of each potential result and to quantify the overall economic impact of postmining land use planning. A discounted cash flow analysis was performed using a surface mine cost model to analyze the sensitivity of selling price to changes in various postmining land use planning parameters. As a result of this analysis, it was determined that the profitability of a project is most sensitive to delays in production, particularly if the delay occurs after major investments have been made. Many methods have been proposed for evaluating postmining land use alternatives (Sweigard and Ramani, 1981). Both quantitative methods, such as benefit-cost analysis (Brock and Brooks, 1968), and nonquantitative methods have been proposed. The application of one specific analytical method is pursued in this paper.

Jan 1, 1988

By Raymond J. P. Brouzes

INTRODUCTION [A Decade After the 1972 Stockholm Conference] The UN-sponsored Stockholm Environmental Conference held just over 10 years ago was the first successful attempt to achieve global recognition of environmental issues. Indicative of the lack of concrete involvement in environmental protection to that time was the fact that only 15 nations had a department or ministry responsible for environmental affairs compared to today's 143. Many governments were uncertain of what were the most serious or pressing problems in their country. There was obviously a need to both identify those issues and to determine how to manage them once identified. At the same time, the public at large was becoming more critically aware that environmental protection had been neglected by government and often ignored by industry. It soon became evident that large amounts of money would have to be spent to redress some of the more evident problems. However, many developing nations were reluctant to slow down the much-needed industrial growth by imposing costly environmental controls on industry. In fact, some waived environmental control altogether as an incentive to attract industry. In some sectors, it was often stated that industrial development and environmental protection were mutually exclusive activities. The need of an awareness, commitment and dedication by all sectors of society to protect their environment was just beginning to develop, replacing the rather simplistic view that all environmental problems were caused by greedy industry.

Jan 1, 1984

By M. A. Marticorena, R. Stratton-Crawley, C. I. Basilio

Diethylenetriamine (DETA) is a selective depressant for nickeliferous pyrrhotite during pentlandite flotation. Laboratory flotation tests conducted on ore and process samples showed that pyrrhotite rejection is greatly improved by small additions of DETA; however, the effectiveness of this depressant is most noticeable when the mineral sample is oxidized. LIMS and XPS analyses conducted on flotation products and pyrrhotite specimens showed that the difficulty in pyrrhotite rejection arises from inadvertent activation of the mineral by heavy metal ions, such as Ni2+, cu2+ and Ag+, that are present in the process water. In the presence of DETA, however, the mineral is deactivated under oxidizing conditions. The deactivation mechanism may involve oxidation of the activation products, which are likely in the form of heavy metal sulfides, followed by solubilization by DETA. FTIR spectra of pyrrhotite electrodes contacted with DETA showed no trace of the reagent on the surface, substantiating the view that its role is one of complexing agent that enhances the dissolution of activation products under oxidizing conditions. FTIR spectra of pyrrhotite contacted with amyl xanthate solutions showed that both dixanthogen and iron xanthate are formed on the surface, the latter becoming more predominant at higher potentials. In the presence of DETA, however, only a small amount of xanthate is adsorbed on the mineral at potentials (Eh) approximately 200 mV higher than the case without DETA.

Jan 1, 1995

By R. Woods, O. A. Kiehn, C. A. Born, F. N. Bender

Development of the flotation reagent scheme at Climax Molybdenum Company's mine at Climax Colorado is reviewed. Features of the mineralogy, milling practice and background of reagent changes are outlined. Special attention is given to the hitherto unpublished details of the application of depressants for copper and lead impurities. The present rougher flotation reagent scheme utilizes a neutral hydrocarbon oil, low in unsaturates and wax, as collector with a sulfated coconut oil emulsifier and pine oil frother. Lime and sodium silicate are used as rougher conditioning and slime dispersant agents. Chalcopyrite and galena impurities are de- pressed in both rougher and cleaner plants by stage addition of Nokes reagent and sodium cyanide to yield a metallurgical grade molybdenite concentrate containing greater than 90% molybdenite and less than 0.5%, 0.1% and 0.05% iron, copper and lead impurities, respectively.

Jan 1, 1976

By Mark S. Dunsbier, Steven J. Schatzel

Large roof outbursts began occurring in November 1986 at Domtar's Sifto Salt Mine in Goderich, Ontario, Canada, working the bedded Salina A-2 Salt unit. The outbursts, measuring about 13-20 m in diameter, released high concentrations of methane from cavities that extended into carbonate roof units. Spontaneous ignitions of methane gas during an outburst seriously burned a mine technician. The Canadian Ministry of Labour contacted the U.S. Bureau of Mines for technical assistance to aid in mine managements1 efforts in alleviating the problem. Initial gas content determinations of salt samples from the outbursts by acoustic test methodology revealed no methane enrichment. In- mine core drilling and subsequent methane content testing indicates the main source of methane in overlying roof units. Drilling the roof units demonstrated the potential of draining methane from areas showing signs of impending outbursting and alleviating the danger. Three mechanisms are considered the most likely means by which methane was introduced to the mine environment. These are (1) hydrocarbon gases released from brines, (2) gas transport along horizontal bedding planes or gas accumulations in bedding separations, and (3) mining-induced fracturing. The abundance of higher hydrocarbons in gas yielded by non-evaporite rock samples from the Goderich mine may be indicative of a former association with liquid petroleum.

Jan 1, 1989

By K. Biswas, H. C. Lee, S. J. Jung

Ore passes are circular openings excavated in rock for in-mine gravity-transportation of ore. It is a very important component of the overall load-haul system of an underground mine. The function of an ore pass is two-fold: first, gravity-transportation of ore from the extraction level to the haulage level and second, the loading operation at the haulage level utilizing the gate assembly. In reality, like any other system component, an ore pass breaks down either due to a hang-up within the ore pass or due to a failure in the gate structure. The failure of the gate assembly may be attributed to a greater extent to the impact load as experienced by the components of a gate structure as a result of the direct impacts of the rock fragments and to a lesser extent to the static or 'dead weight' loads of the accumulated materials in the ore pass. This paper uses simulations in Working Model® as an aid to investigate the effect of both the impact and static loading on the gate assembly and on the inner-walls of the ore pass. The paper also investigates the impact of ore pass inclination and the presence or absence of dogleg on the static and impact loading of the same.

Jan 1, 2002

By A. P. Schissler, S. Badr, U. Ozbay, M. D. G. Salamon

Yield pillars are used in deep coal longwall mines, where the depth of cover is so great that a conventional chain pillar design tends to be unsuccessful. The design of yield pillars in deep U.S. mines has been based mainly on experience, pressure arch theories, numerical modeling, and empirical methods that originated in the eastern U.S. coal mines. This paper offers field experience based guidelines for the design of yield pillars through analysis and evaluation of numerous case histories. Case histories of yield pillar applications were collated and critically reviewed to identify trends in data that can be used to formulate the design guidelines. Convincing trends observed in the case of successful case histories indicate that yield pillars should have a width to height ratio in the range from 2.8 to 6.0, a stability factor during development between 0.4 and 0.6, and shed between 94% to 98% of the cover and abutment load, as provided by the Carr-Wilson (1982) method. Also, yield pillars should be utilized in coalmines only where the roof and floor rock have a CMRR rating of 50 or better.

Jan 1, 2003

By A. M. Al Taweel, V. K. Kasireddy

Intensive coal cleaning usually necessitates fine grinding of the feed in order to liberate inorganic constituents from the surrounding organic matrix. This however adversely affects the ability of flotation techniques to beneficiate coal. Such difficulties are mainly attributed to the increase of particle population density, the decrease in capture efficiency, and the greater tendency of inorganic components to report with the froth. A three-pronged approach was adopted in an attempt to overcome these difficulties. The aggregative flotation technique was used to reduce particle population densities and improve capture, while unconventional spargers were used to generate a wide range of bubble sizes. Use of unagitated flotation columns allowed for the maintenance of countercurrent flows in both the recovery and washing sections. The combined effect of these parameters enabled the recovery of 92% of the carbon content of coal having an average particle size of 20.5 µ m, while rejecting 85% of the ash. This was accomplished using 2 kg of fuel oil per tonne of coal in a column 2 meters high operating at 2-3 tph/m3, a capacity which compares favourably with conventional cells.

Jan 1, 1989

By Gregg Sherry, Bob Gordan

INTRODUCTION This paper report presents the general case history for the construction of the North Outfall Replacement Sewer (NORS) project. A 1982 City of Los Angeles Wastewater Facilities Plan recognized that a 19.3 km (12¬mile) section of the North Outfall Sewer (NOS), from Culver City to the Hyperion Treatment Plant, was deteriorating and had insufficient capacity. Built in the 1920s of unreinforced concrete and lined with vitreous clay tiles, the originally unventilated NOS suffered considerable corrosion by H2S gas. This caused roof tile and concrete to fall to the bottom of the pipe, which, over the years, reduced the 11.9 M3/S (420 cubic feet per second (cfs)) design capacity of the NOS to a current level of approximately 8.5 M3/S (300 cfs). This worsening condition, thereby contributing to the polluting of Santa Monica Bay. This situation culminated in the State of California Regional Water Quality Control Board, Los Angeles Region, issuing a Cease and Desist Order on January 27, 1986, which required the City of Los Angeles to eliminate such dry weather overflows through a number of prescribed actions. One of these was to construct the North Outfall Replacement Sewer (NORS) by April 1, 1993. PROJECT DESCRIPTION The NORS Project involved the construction of over eight miles of tunnel and related structures. It consisted of a main tunnel conduit to convey sewage to the Hyperion Treatment Plant and four diversion lines connecting the main tunnel to major existing

Jan 1, 1993

By W.A. van der Westhuizen

Mineral resource management has become a well used phrase in the mining and minerals industry. Mineral resource managers are being appointed in many mining and mining-related organisations. Every mineral resource manager has his/her own ideas regarding what MRM means. There seems to be a tendency to define MRM close to home; geology would apply MRM principles to their environment and so will metallurgy, planning, mining, survey and every discipline and participant in the mining value chain. What do we do in the MRM environment and what do we aim to achieve? Surely we must have goals: what are they? Is value tracking MRM? What about best practices? Let us start at the beginning: what do we want to achieve in any mining operation? The minerals industry would like to make money, now and in the future. This calls for serious management inputs. The most important aspect is a holistic management approach regarding the total mining value chain. Process optimisation follows with application of Mineral Resource Throughput Management (MRTM) and Process Flow Management principles. Once flow processes have been established the necessary condition driven standards and measurements will be installed to ensure continuity in applicable processes.

Jan 1, 2012

By J. Peshel

This paper presents a regulatory perspective of the design and construction of the Exploratory Shaft Facility which may become part of the future high-level waste repository. Site characterization concerns and long-term waste isolation concerns require the consideration of various shaft design and construction aspects quite different from that of conventional shaft sinking.

Jan 1, 1989

By P. D. Roa, O. Noirot, H. Owens, D. E. Walsh, J. R. Mokka

Over a period of eight years (1989 through 1997) the Mineral Industry Research Laboratory (MIRL) at the University of Alaska, Fairbanks, AK, has investigated the effects of temperature, residence time and particle size on the degree of low-rank coal (LRC) upgrading for a nonevaporative, hydrothermal drying process, also known as hot water drying (HWD). The following three Alaskan LRCs were used in this study: Usibelli coal (Seam No. 4) from the Nenana Coal Field, Little Tonzona coal from near McGrath, and Beluga coal from the Beluga-Yentna field on the west side of Cook Inlet. Replicated factorial tests were conducted. Experimental results indicate that most of the low-rank coal upgrading via HWD occurs rapidly within the first 10 to 20 min of residence time. Thereafter, the upgrading process slows down, and more than 100 min of additional residence time is required to produce percentage increases equivalent to what was achieved in the first 10 to 20 min. The properties of the hot water dried products were very sensitive to process temperature. Calorific value, carbon and oxygen contents, equilibrium moisture levels and Hardgrove gindabilities all appear to change linearly with temperature within the 275°C to 325°C (525°F to 61 5 OF) range.

Jan 1, 2000

By Scott Jones, Tony R. A. Durkac, David W. Stephens

Williams Mine was considering spending a large amount of capital to install a pumping system underground to rid the mine of slimes; with the introduction of Delvo Stabilizer this challenge would be met. This paper will address the process to introducing the Delvo Technology into the Cemented Roddill Program. Some areas that were worked on included: testing of the mix design, bet dosage rate for an economical solution, temperatures at different levels of the mine, the amount of time needed once the grout was put to sleep, preventing plugs in the delivery system A controlled dosage that gave the Operator a 12 hour window of flowable material was developed and an added feature to the introduction of Delvo was also noted; an increase in MPa strength over the design. Collectively with the help of operators at the batch plant and the underground miners involved with the program all concerns where addressed and all goals met. A cost effective program was implemented thus increasing productivity, preventing costly plugs, and improved the actual maintenance on the entire delivery system, which included the batch tanks, aggitator tanks, all pumps required in the system and of course reducing buildup of slimes at the cones.

Jan 1, 2001