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By N. C. Miller

In situ stope leaching is an innovative mining system that reduces the surface impacts of conventional underground mining. As in any leaching operation, stope leaching requires hydrologic site characterization and control of leaching solutions. Several tracer tests, using sodium chloride (NaCl) in various concentrations, were conducted by the U. S. Bureau of Mines (USBM) in an underground experimental stope to simulate a leaching environment. The tests were conducted in fractured crystalline rock that was partially saturated. As tests progressed, the surrounding fractures became more saturated. A predominant flow direction away from the stope was found during testing. Testing procedures, preliminary results, and remediation ideas based on field observations are presented in this paper.

Jan 1, 1995

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 Charles R. Nelson

INTRODUCTION The support of openings in weak sandstones can be achieved by spraying on a liquid grout which will soak in and harden to form a shell. In the St. Peter sandstone of the Minneapolis-St. Paul area of Minnesota compressive strength is increased from about 0.7 MPa (100 psi) to over 7 MPa (1000 psi) to depths of 152 mm (6 in.). A shell of this strength and thickness can pro¬vide support much as shotcrete does but at lower cost. If the spray-grouted shell is inadequate for the support needed, it provides a good base for shotcrete application, which is often difficult in weak or friable sand¬stones because the fresh shotcrete peels off along with a thin layer of sandstone. The spray-grouted technique was initially developed at the Civil and Mineral Engineering Dept. of the University of Minnesota using funds provided by RANN Div. of the National Science Foundation. It was used as temporary support in a 3-m (10-ft) wide, 3700-m (12,000-ft) long storm water drain tunnel by the Minneapolis Sewer Construction Dept. Much of the field development at this construction project was carried out with the cooperation of the personnel of the city of Minneapolis. It also was used for temporary support in a 3.3-m (11-ft) wide, 1800-m (6000-ft) long storm water tunnel built in 1978-1980 for the Minnesota Dept. of Transportation in Minneapolis. In 1975, 2000 m (6500 ft) of 1.5-m (5-ft) wide utility tunnels were spray-grouted for final lining. A 50-m (150-ft) long, 1.2-m (4-ft) wide test tunnel built at the university in 1974 was sprayed for half its length and is being observed for long-term behavior which to date (1981) is good. GROUT MATERIALS The requirements for the liquid grout are more severe than those for injection grouting. Besides being able to penetrate the rock and develop sufficient strength, it must have the following properties: (1) be nontoxic, noncombustible, and have a low odor for spraying under¬ground; (2) have controllable viscosity and setting time for adjusting to the permeability of the sandstone and the required depth of penetration; and (3) must "wet" the sandstone and develop capillary "draw" to penetrate to the required depth. A sodium silicate-based grout with the proper setting agent meets these requirements in the local St. Peter sandstone which has a uniform grain size dis¬tribution curve with D,,, = 0.1 mm, permeability of about 10 to 20 darcys, a porosity of about 25%, and less than 5% of silt size or smaller. It is 97 to 99% pure SiO2. The grout mix consists of Philadelphia Quartz Co. Type N sodium silicate and Celtite 55 Terraset Spray Grade (SG) distributed by Celtite Inc., Cleveland, OH. A volume mix ratio of 100:9:125 for sodium silicate to setting agent to water produces a grout with two to four centipoise viscosity, and initial set at 20°C of 20 min. Setting time is temperature dependent, being longer at lower temperatures, but can be adjusted by varying the setting agent concentration a few percentage points. APPLICATION The application is simple. Spray the grout on the surface until the desired penetration is achieved. Penetration rates of about 5 mm (0.2 in.) per min. are realized in the St. Peter sandstone for up to 152 mm (6 in.) of penetration (30 min. of spraying). The set time must be longer than the spraying time. The spray nozzle should be moved back and forth at a rate that minimizes runoff due to surface buildup. Both a single large hand-propelled nozzle and multiple small machine propelled nozzles have been used successfully. The pumping and mixing equipment used consisted of: (1) a hand-held, hand-pumped, 11-L (3-gal) gar¬den sprayer for test patches (mix by shaking); (2) 76-L (20 gal) batch mixing in barrels with small electric pumps for spraying and mixing; (3) two component pumps consisting of Hypro Model 5300 piston pumps (Hypro Pump Co., St. Paul, MN) on a common shaft driven by an electric motor. The sodium silicate is premixed with a portion of the water and the setting agent is mixed with the rest of the water. The two components are combined and mixed just before the nozzle; and (4) a proportioner metering the setting agent just before the nozzle with the water and sodium silicate premixed and pumped from the surface. The proportioner permits the bulk of the grout to be pumped through a single pipe or hose with only the setting agent stored underground. This is efficient and low in cost. COST The grout mix costs about $0.20/L ($0.75 per gal) in 1978 dollars. Filling all the voids 100 mm (4 in.) deep (25% initial porosity) would have a materials cost of about $5.40/m2 ($0.50 per sq ft). The labor cost of application is usually less than the material cost for normal tunnel jobs. A pump and proportioner cost less than $1000. Sodium silicate storage tanks and pip¬ing costs would depend on the particular site conditions. REFERENCES AND BIBLIOGRAPHY Nelson, C. R., 1977, "Spray Grouting for Tunnel Support and Lining," Underground Space, Vol. 1, No. 3, pp. 241¬ 245. Yardley, D. H., Nelson, C. R., Stocker, T. H., 1974, "So¬dium Silicate Spray Impregnation of Tunnels in the St. Peter Sandstone," Research Report, University of Min¬nesota, 118 pp.

Jan 1, 1982

By Yi Luo

Subsidence events occurring above inactive room-and-pillar (R&P) mines, including abandoned mines or sealed portions of active mines, are classified as unplanned subsidence, because of their timing and distribution are unpredictable. When investigating such subsidence cases, many uncertainties exist and disputes often focused on whether the observed structural damages are caused by mine subsidence or other natural or artificial causes. However, solid evidence to identify the cause is often lacking. In order to conduct a complete and impartial investigation, thorough knowledge of subsidence theories and other ground control topics, as well as broad experience, are required.

Jan 1, 2011

By J. L. Parks

Designs, installation procedures, and performance characteristics of martensitic cast steel concaves are reviewed. Longer wear life, fewer pieces to install, and eliminating the need to scarf joints are the principal benefits of using martensitic steel to replace austenitic 12% Mn (Hadfield's) steel, the traditional concave material for large gyratory crushers. Installation procedures that reduce the risk of breakage are described. Operating practices that affect concave performance are also discussed.

Jan 1, 1992

By Matthew Wilcock, Kenichi Soga, Peter Wright

"This paper discusses how an existing grey cast iron segmental tunnel lining responds to construction of a new SCL tunnel transversely beneath with zero clearance; how the damage assessment for the existing tunnel was conducted for impact of longitudinal movement; how the tunnel is subsequently instrumented to observe both movement and stress manifestation at its longitudinal connections and how the results from the instruments compare with the prediction. An assessment of neutral axis of the tunnel in longitudinal flexure is also explored using interpolated horizontal displacement data from a conventional monitoring system. INTRODUCTION The Royal Mail Tunnel (RMT) (formely Post Office Tunnel or ‘Mailrail’) was constructed using semi-mechanised deep level construction techniques from 1914 and 1917 (Lee 1938). Commencing full commercial operation in 1926, the system was until 2003; conveying parcels the 6.5 miles between seven sorting offices in the center of London, and linking the system with main line rail stations, Paddington and Liverpool Street (The British Postal Museum & Archive 2013). A single tunnel of 2.74m internal diameter facilitated a bi-directional twin narrow-gauge (~611mm gauge) railway road system for electric driverless operation. The tunnel lining is segmental grey cast iron (grade 10) made up of six flanged segments and a smaller flanged key segment; constructed into a ring of width 508mm The external diameter of the tunnel is 2.971m. For further details of the tunnel lining and material properties are outlined by (Devriendt & Alhaddad 2014).London Underground Bond Street Station is a major station of the London Underground system in the very center of London; in its vicinity, the RMT runs west-east perpendicularly above the north south Jubilee Line tunnels. Historic survey record of the Jubilee Line’s (the then Fleet Line) construction in 1974 and its impact on the RMT case is documented in (Crossrail 1993): it is one of only a hand full of case-studies frequently cited when assessing the impact of new construction scenarios of this type; not least including the assessment of the same tunnel, at the same site some 40 years later, and the subject material for this paper: It noted significant vertical displacement of ~90mm to the RMT without noticeable serviceability or operational impediment."

Jan 1, 2016

By P. Zhang, W. D. Xiao

"A process mineralogy study was conducted to guide the development of a process for the extraction of vanadic oxide (V2O5) and aluminum compounds from bone coal and its ash as well as the production of byproducts. Results show that to extract the targeted metals at higher recovery rates, one must find a suitable agent capable of destroying the structure of the mica and breaking down the lattices in which these metal-bearing minerals occur. For such minerals, metal extraction efficiency is usually low if the approach is simply adding an oxidant that cannot destroy the structure. In this research, a byproduct fluosilicic acid (H2SiF6) at 20 percent concentration from an operating phosphate fertilizer plant was found to be powerful and effective for the leaching of vanadium-bearing minerals such as mica. This environmentally friendly process improved the leachability of valuable constituents in the bone coal significantly, with the leaching rate of V2O5 increased to 94 percent, from 83 percent, and that of aluminum oxide (Al2O3) increased to 75 percent, from 40 percent. X-ray diffraction and chemical analyses showed that the structure of mica can be destroyed using fluosilicic acid either alone or in combination with sulfuric acid. Based on the process mineralogy study, a processing flowsheet was tested, which resulted in a final V2O5 powder product with purity of 99.2 percent and an overall recovery rate of higher than 85 percent. The flowsheet consists of leaching, ion exchange and ammonium precipitation. With this process, various byproducts could also be generated, including aluminum fluoride, cryolite, aluminum sulfate, calcium sulfate, ammonium chloride, sodium chloride and cement admixtures.IntroductionVanadium is an important rare metal broadly used in the production of steel alloys and nonferrous alloys as well as in the chemical field as a catalyst. Vanadium-bearing bone coal and clays are abundant in China, and they contain other valuable metals such as aluminum. Bone coal is a type of carbonaceous shale. The vanadium-bearing minerals mainly exist in mica, kaolin, iron oxides, tourmaline and garnet. According to previous studies (Xiao, 2007; Xu, 1989; Xu and Chen, 1991; Ji and Xu, 1991; Chen and Xu, 1992; Xu and Heng, 1992), vanadium occurs from the isomorphous substitution of V3+ for Al3+ or Fe3 + in the crystal lattices of mica, tourmaline, garnet, dolomite and aluminosilicate minerals. In iron oxide and kaolin minerals, vanadium appears as an adsorption form. The isomorphous substitution of vanadium is more stable in tourmaline, garnet and silicate minerals. Therefore, vanadium in bone coal may exist in three phases: vanadium-containing kaolin and oxides, vanadium-containing mica, and tourmaline and garnet (Xu and Chen, 1993; Xu and Zhang, 1993; Xu and Wang, 1994; Feng, et al., 2005; Lin and Fan, 2001)."

Jan 1, 2015

By J. E. Golsan

Barrick Gold Corporation installed an expert style control system on the grinding circuit at its Mercur mine. Barrick's main objective was to increase mill throughput by a minimum of 3%. The expert system team collected operating data and then created a knowledge base using Comdale/X expert system development software. The expert system moved from on-line "advisory" mode, testing of expert rules, to full process control in four weeks. Expert system control benefits include a 9-15% increase in mill throughput, consistent mill loading leading to less wear on SAG mill parts, and improved downstream process control.

Jan 1, 1998

By L. M. Sandoval, R. Kaunda, C. N. Vasquez, S. C. Falcon

Artisanal and small-scale mining (ASM) involves a wide range of practices, from highly informal and minimally mechanized mining operations with predominantly manual work to formal practices carried out by small but high-tech industrial mining operations [1]. ASM may be a traditional practice, with skills passed from one generation to another. It can be an important driver of local economic development, while simultaneously contributing to a complex array of positive and negative socioeconomic and environmental impacts. There are more than 100 million artisanal and small-scale miners in 80 countries worldwide, and they supply around 20% of the worldwide gold production [2]. According to the Peruvian Ministry of Energy and Mining (MINEM), in 2020, 32% of fatal accidents in formal (legal) Peruvian mining operations were caused by groundfall events [3].The lack of resources, knowledge of best practices in health and safety, and insufficient equipment used in ASM make the probability of suffering work accidents very high. Furthermore, groundfall is one of the main causes of accidents and injuries in underground mines around the world [4], and it is also a concerning problem in ASM, where the associated fatality rate is similarly high. Unfortunately, a proper record of accidents in ASM is not kept, so an estimate of the actual number of groundfall accidents cannot be accurately developed [5]. Currently miners rely on empirical and observational tools acquired through previous experience to assess groundfall risks. The risks for artisanal miners can tend to be significantly greater than risks for those from the mining industry. This is due to the lack of regulation of ASM, which tends to operate outside the legal framework and the regulations lack of enforcement in force for occupational health and safety regulations.

Mar 2, 2022

By Gabino Preciado, Louis Sandbak

Corrosion has long been identified as a potential risk to the long-term ground support stability due to degradation of internal ground support elements such as rock bolts but also to surface support from mesh, straps, bolt plates, and ancillary support of utilities such as hangers. The emphasis is to understand the corrosion mechanisms on the primary ground support elements and any supplemental steel or arch sets. The water and chemical processes from oxidation(rust), stress corrosion cracking, and even the electrical charge created during these reactions contribute to the corrosion of the steel support elements. The main goals are to mitigate the effects of corrosion and to be able to predict those areas of high water or high sulphide for potential rehab maps. It is also essential to identify and classify the visible corrosion as part of damage mapping and provide bolt testing to those areas of suspected internal damage from corrosion. This is important to get an estimate of the support longevity or expected life before rehab to minimize the overall cost of the operation.

Feb 1, 2024

By D. Blackwell, W. R. Reed, G. Ross, S. Peters, M. Shahan

Dust collectors for roof bolting machines generally use a dry box to collect the roof bolting material. Recently, an underground mining operation converted a dry box dust collector to a wet box dust collector with a unique exception from MSHA for testing purposes. Water is routed to the roof bolter from the main water line of the continuous miner. The wet box utilizes a water spray to wet the incoming material. Testing was conducted comparing the two different collector types. Respirable dust concentrations surrounding the roof bolter with the different collection boxes were similar. The main difference in respirable dust concentrations occurred when cleaning the dust boxes. The average respirable dust concentration during cleaning of the wet box was 0.475 mg/m3, and during the cleaning of the dry box the average respirable dust concentration was 1.188 mg/m3. The silica content of the roof material was high, ranging from 28.9% to 52.7% during this study. The results from this study indicate that using the wet box as a collector reduced exposure to respirable dust up to 60%when cleaning the collector boxes. INTRODUCTION Currently, the dry vacuum dust collection system is the standard dust collector on roof bolters in underground coal mining. It has been proven to be an effective dust control to prevent roof bolter operators’ overexposure to coal mine respirable silica dust during roof bolting operations [Potts eta l, 2011 and Colinet, 2013]. The dry collection system works as follows: drill cuttings are vacuumed into the precleaner. The precleaner dumps oversized material onto the mine floor while undersized material enters the dust box via vacuum airflow produced by a blower. The cuttings entering the dust box are either deposited on the floor of the dust box or into a dust collector bag. The vacuum airflow leaves the dust box through a final filter which removes any small particles from the airflow. Figure 1 shows the internal view of two dry collection system boxes; one that does not use the dust collector bags and another that uses the dust collector bags. Dust collectors that use dust collector bags to contain material inside the dust collector box have been shown to provide better protection to roof bolter operators than dust collectors that do not use the bags [Listak and Beck 2008]. The dry collectors are very efficient at reducing respirable coal mine dust exposure from roof bolting activities. Recently, Blue Mountain Energy’s Deserado underground coal mining operation, outside Rangely, CO, made a modification to the dry collector system to convert it to a wet collector system with a unique exception from the Mine Safety and Health Administration (MSHA) for testing purposes. The wet collector system still uses a vacuum system to pull material from the vertical roof bolt hole, but incorporates a water spray nozzle located inside the dust box to wet the incoming material. The flow through the nozzle ranges from 1.9–7.6 lpm (0.5–2.0 gpm) at 0.69 MPa (100 psi). The conversion to the wet collector system removes the precleaner and the cyclones inside the dust box. The wet material (sludge) then empties through a specialized valve in the bottom of the box. The final filter is still used to prevent respirable dust from leaving the dust box through the exhaust. During operation, the wet collector dust box is opened and hosed clean after roof bolting every place. After every two entries the final filter is changed out. Figure 2 shows the inside of the wet collector box.

Jan 1, 2019

By William Gleason

While a fleet of 100 t (110 st) haul trucks were busy removing overburden from the open pit mine in the distance, a host of dignitaries, including congressmen and California state representatives, gathered for Molycorp?s official groundbreaking ceremony at the Mountain Pass Mine on April 20. The mine, recognized as arguably one of the richest rare earths deposits in the world, is at the center of Molycorps? strategic ?Mine to Magnets? plan. It aims to secure America?s future supply of rare-earth-dependent technologies by developing a strong domestic rare earth industry. Molycorp identifies rare-earth-dependent technologies as those that cannot be produced without rare earth materials. The list of these technologies is long and growing. It encompasses everything from personal electronics and batteries for hybrid cars to permanent magnet generator wind turbines and national defense applications.

Jan 1, 2011

By V. Nasri, S. Rashidi

The Baltimore Red Line Project is a proposed 22.7-km (14.1-mile) long east-west light rail transit (LRT) line envisioned to connect the areas of Woodlawn, Edmondson Village, West Baltimore, downtown Baltimore, Inner Harbor East, Fells Point, Canton and the Johns Hopkins Bayview Medical Center Campus. The Red Line LRT System has two tunnel segments - the Cooks Lane Tunnel (CLT) and the Downtown Tunnel (DTT). The CLT segment is roughly 1,920-m (6,300-ft) long. It commences at the west portal located at the highway ramp for I 70 (to be removed) and terminates at the east portal that is at the intersection of Edmondson Avenue (U.S. Route 40) and Glen Allen Drive. This segment of the project consists of the following construction components: approximately 1,458 m (4,786 ft) of tunnels, 143 m (469 ft) of cut-and-cover tunnel, and 318 m (1,045 ft) of retained cut (U) section. The approximate horizontal alignment for the Red Line LRT Project is shown in Fig. 1.

Jan 1, 2014

By J. Baz-Dresch

A custom-designed shotcrete system, consisting of an 8-m3 (10-yd3) transit mixer and a shotcrete spray boom carrier, is used bya 2-man crew to transport and place up to 25 m3 (32 yd3) of steel-fiber reinforced, dry, silica fume shotcrete per II-hour shift. Shotcrete is custom blended by a local supplier and delivered predampened. Mix design and quality control is performed with a computer program that develops and graphs a properly balanced mix based upon the materials sieve analyses.

Jan 1, 1991

By B. P. Atkin, D. W. Brown

The handleability monitor has been specifically designed to provide a simple, rapid and on-site assessment of the whole or part consignment of a coal blend in its ‘as-received’ condition. Evaluation has been carried out at a number of coal-handling plants, on both metallurgical and power station coals from a variety of sources world-wide. The results which are used to predict which coals could routinely pass through a handling plant at line capacity, without causing blockage or delay, have also been correlated with moisture, ash and size analyses of the coal to develop a predictive model of handleability.

Jan 1, 2000

By N. Kuyucak

While Acid Rock Drainage (ARD) occurs naturally, its management poses challenges especially at the decommissioned sites. Two passive systems were designed and implemented at two different decommissioned mining sites in Canada in order to treat the seepage with ARD characteristics. One seep originating from a reclaimed tailings pond was neutral mine drainage (NMD) containing circumneutral pH and dissolved metals (e.g., nickel, zinc) in elevated concentrations. The other ARD from a tailings stack showed net acid drainage characteristics with low pH and higher metal concentrations than regulated limits. Due to the different ARD characteristics occurring at each site, different types of treatment systems were designed and implemented. Both sites experience extreme winter conditions. The NMD has been treated mainly by limestone drain and aerobic systems since 1997. The acid ARD has been treated by a system consisting of an anaerobic sulfate reducing bacteria (SRB) cell, an aerobic cell and a limestone filter since 2004. As this site has no access roads, space, or power, a significant capital cost as well as sludge disposal/storage requirements would have been required to institute a chemical (active) treatment system. Both systems have been successfully treating seepage and producing a water quality year round that could comply with regulated objectives.

Jan 1, 2010

By C. Vass, A. Noble

"The chemical treatment of mine-influenced waters is a longstanding environmental challenge for many coal operators, particularly in Central Appalachia. Mining conditions in this region present several unique obstacles to meeting NPDES effluent limits. Outlets are often located in remote areas with challenging terrain where conditions do not facilitate the implementation of large-scale commercial treatment systems. Furthermore, maintenance of these systems is often laborious, expensive, and time consuming. Many large mining complexes discharge water from numerous outlets, while using environmental technicians to assess the water quality and treatment regime multiple times per day. As an alternative solution, this paper describes the ongoing research and development of automated chemical treatment systems for mine water discharges. The current work explains the development of a bench-scale prototype that has been used to evaluate various monitoring, notification, and control strategies. Results from this testing are compared to present day practices to show the costs and benefits of implementing an automated treatment and monitoring program. INTRODUCTION Background The National Pollution Discharge Elimination System (NPDES), per Section 402 of the Clean Water Act, requires permitting of all point source discharges from coal mines. This program dictates that all water leaving a permitted boundary, whether it be process water or storm water, must leave through a monitored NPDES outlet. These permitted NPDES discharge points, also denoted outlets, outfalls or monitoring points, must meet a rigorous set of effluent limits to maintain compliance. Common effluent limits monitored by coal companies include but are not limited to iron, manganese, aluminum, total suspended solids (TSS), and pH. Raw mine water is often deficient in meeting these effluent limits, and as a result, various treatment processes are required to bring the water into compliance prior to environmental release. Depending on the mining and geologic conditions, water monitoring and treatment programs can constitute a significant operating cost for the mine operator. In Appalachia, acid mine drainage (AMD) constitutes one of the most significant and widespread water quality challenges that must be treated per NPDES guidelines. AMD is a general lowering of water pH that occurs when water and sulfide minerals interact within oxidizing conditions in coal and metal mining. Acid Rock Drainage (ARD) is a similar term that applies more generally to non-mining causes such as highway construction and other large-scale excavations. Today, mine influenced waters (MIW) is becoming more popular as the general term referring to all types of mining-induced water issues. AMD has been researched substantially over the last 30 years, with investigators studying various causal mechanisms and treatment strategies (e.g. Gazea, 1996; Ziemkiewicz, 1998; Skousen, 2000; Kleinmann, 2001; Johnson, 2003; Akcil, 2006; Kalin, 2006). This research has produced several pragmatic results, which provide numerous methods to predict, prevent, and treat AMD."

Jan 1, 2016

By B. P. Atkin, D. W. Brown

A handleability monitor was designed to provide a simple, rapid, on-site assessment of the whole or part consignment of a coal blend in its "as-received" condition. The monitor was evaluated at a number of coalhandling plants, which handle both metallurgical and power station coals from a variety of sources worldwide. The results can be used to predict which coals could routinely pass through a handling plant at line capacity without causing blockage or delay. Despite extensive analysis of the data, it was not possible to correlate the handleability index to the regular sample analysis of moisture, ash and size analyses of the coal blends.

Jan 1, 2002

By Ihsan Berk Tulu, Ted Klemetti, Deniz Tuncay

Accurately estimating load distributions and ground responses around underground openings plays a significant role in the safety of the operations in underground mines. Adequately designing pillars and other support measures relies highly on the accurate assessment of the mining-induced loads, as well as the load-bearing capacities of the supports. There are various methods that can be used to approximate mining-induced loads in stratified rock masses, both empirical and numerical. In this study, the numerical modeling approach recently developed at West Virginia University, which is based on the modeling approach developed by the National Institute for Occupational Safety and Health (NIOSH), is investigated using the finite difference software FLAC3D. The model includes the longwall panels, the adjacent chain pillar systems, and the different stratigraphic layers of the overburden. Using the 3D model, changes in loading conditions and deformations on the areas of interest, induced by an approaching longwall face, can be examined. This paper details the 3D modeling of a longwall panel utilizing this approach, and the verification of the results against field observations. The studied panel was 360 m wide with a 3-entry chain pillar system and about a 160-m average overburden depth around the studied area. The overburden strata consist of alternating layers of shale, sandstone, and limestone. The FLAC3D results were compared against field measurements from the mine site. The stress change values measured in the chain pillars were comparable with the modeling results. The model also replicated the surface subsidence profile obtained from field measurements fairly well. Overall, the 3D modeling approach was found to be successful for the case study longwall panel.

Sep 23, 2020

By D. R. McMullin

The purpose of this paper is to provide a method for mine planners and engineers to evaluate on a conceptual basis if a shaft or a ramp is most suited for their particular underground access requirements. Mine planners and engineers are often raced with the decision of choosing which method of underground access is best suited for their particular requirement. Underground access can be provided by a shaft, ramp, adit, or slope for either exploration and/or actual production operations. This paper will focus on the construction cost and schedule differences between sinking a shaft and excavating a ramp which are currently the most typical access alternatives used in the western United States. It would be wrong to state that one method is superior to the other in every case. There are many variables to be addressed, and the degree of sensitivity of each of them will eventually dictate which method is the most suited to the particular circumstances. This paper compares the construction cost and schedule or sinking a shaft to excavating a ramp at different depths given different ground and hydrological conditions. This comparison provides mine planners with "rules or thumb" that may be applied to the access selection process.

Jan 1, 1994