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By D. Armour, E. R. Bauer, D. R. Zuchelli, T. M. Durr, E. R. Reeves

A team of noise specialists from the National Institute for Occupational Safety and Health (NIOSH), Mine Safety and Health Administration (MSHA), and Joy Manufacturing attempted to evaluate an engineering noise control on a Joy longwall stageloader. The team developed and executed a test plan to evaluate the noise attenuation of the engineering control in an underground coal-mining environment. An evaluation of the long-term noise effects and durability of the engineering control were also a focus of the research effort. Pre- and postcontrol measurements consisting of a combination of worker dosimetry and task observations, tape recordings, and sound level measurements were used to evaluate the engineering control. A description of the engineering noise control, pre- and post-control sound level measurements, and an evaluation of control success is presented. In the end, no evaluation was possible due to mining practices, miner task variations, and inconclusive data.

Jan 1, 2005

By F. H. Muge

This paper intends to present a general methodology implemented in a software package and which can be applied, whatever the exploitation methods, mining constraints and time basis, to mine planning and production schedulling. Examples are: presented covering the medium-term and short-term planning procedure and the production schedulling for two underground mines located in the South of Portugal (Neves- -Corvo and Aljustrel Mines). Both are complex suiphide orebodies with differ- ent exploitation methods (Neves-Corvo: cut-and-fill; Aljustrel: Long-hole stop- ping).

Jan 1, 1992

By Nicola Della Valle

The Singapore Land Transport Authority (LTA) is currently expanding capacity by extending their Metro system from South of the city to the North East end of the island. The scope of Contract 704 involved excavating a long reach of tunnel which had a complex rock/soil interface comprised of fresh granite overlaid by clayey sand. Two 6.5 m diameter Earth Pressure Balance (EPB) TBM’s were used to bore the tunnels. These machines were designed so that both hard and soft rock could be excavated. The cutter head was equipped with a combination of both rippers and disk cutters. Soil conditioning was achieved by delivering foam through a multiple port injection system. This very irregular profile of the interfaces posed many challenges to the equipment and the people.

Jan 1, 2001

By Shahzad Moosa, John M. Pitt

Statistical variability in rock strength can be evaluated by rapid, inexpensive ultrasonic pulse velocity measurements, an important consideration since rational probabilistic design in rock is often hampered by the cost of obtaining sufficient strength data. Experimental ultrasonic pulse velocity measurements on a set of 96 limestone specimens showed two distinct statistical populations. Both populations followed Weibull distributions, with flexural strength distributions paralleling pulse velocity distributions. Thus, experimental results suggest that strength variability may be predicted from a few rationally selected strength measurements and several ultrasonic tests. The correspondence between strength and velocity may then be used to develop an experimental transformation function, allowing the development of a statistical distribution function for strength based on the known statistics of velocity. Comparison of experimental groups with full sets of corresponding strength1 velocity data showed the technique offers a fairly accurate description of the statistics of strength.

Jan 1, 1986

By A. Gene Collins

The demand for lithium may increase because of its potential use id lithium batteries and in the generation of electrical power by fusion. Therefore, lithium abundance in some United States oil-field brines was surveyed. Oil-field brines from numerous locations were found to contain more than 100 mg/l lithium and several exceeded 500 mg/l. An area of 25,000 km2, with a reservoir thickness of 60 m, an effective porosity of 5 percent, and a brine with a specific gravity of 1.2 containing 100 mg/l of lithium, contains about 0.75 x 106 tonnes (1 tonne = 1 metric ton = 1,000 kg) of lithium in 7.5 km3 of brine. It is postulated that the brines were enriched in lithium by lithium-rich spring water or lithium leached from volcanic rocks followed by evaporation sequences.

Jan 1, 1978

By M. Kalin, H. D. Sharma

INTRODUCTION The inactive uranium mill tailings investigated in this study are located in two mining districts, Elliot Lake and Bancroft, Ontario, Canada. The sites exhibit a mixture of surface features consisting of dry areas, with or without vegetation and areas covered with water. On the edges of the water bodies, indigenous vegetation has invaded the tailings beaches [ Typha] spp. (Tourn.) L. a dominant plant on these tailings beaches, has been studied for the uptake of radium-226 and lead-210 (Kalin and Sharma, 1981b) from the tailings. It was found that most of the radium-226 remains in the roots of the plants, and that the solubility of radium-226 in control soil differs from that in tailings. The uptake of radium226 by vegetation and other biota is related to the solubility of the element in water. Factors controlling the solubility of radium-226 in uranium mill tailings are of interest in assessing the environments effects of these wastes. Rusanova (1962) found that the soluble or extractable amount of radium from the soil is inversely related to the total concentration of calcium and magnesium in the soil. Experimental work has clarified some aspects of the leachability of radium from uranium mill tailings (Levins, et al., 1978; Wiles, 1978, and others). Halvik, et al., (1967) studied the effects of pH and chemical composition of surface water on the liberation of radium from uranium mill tailings and uranium ore. He found that an increase in pH, up to a value of pH 9, decreases the amount of radium released from the tailings and the ore. A positive effect of calcium chloride was noted on the leachability of radium. Benes (1981) reviewed the physicochemical forms of radium and its migration in water. Based on experimental work, he identifies primary factors which determine leaching of radium from uranium mill tailings. The ratio of the volume of the leaching solution to the weight of the leached sample; the composition of the leached solids and the leaching solution, and finally the pH of the leaching mixture are of importance. He emphasized that a scarcity of field data exists, which would relate experimental work to the actual situation in the tailings ponds. Inactive tailings ponds in Ontario are 16 to 23 years old, and processes which are of importance in evaluating the long term effects of uranium mill tailings in the environment can be studied. The objective of this work was to investigate the leachability of radium-226 from the tailings under field conditions. MATERIALS AND METHODS Sample Collection A description of inactive tailings sites in Ontario, where the tailings and the water for this study were collected, can be found in a report by Kalin (1981). [Typha] spp. specimens were excavated from 16 different locations along with the tailings attached to the roots of the plants. A tailings sample from the surface area, around the plant, was secured before the excavation. Tailings samples were also collected from dry areas free of vegetation in the vicinity of the wetland stand. Surface samples (depth 0-5 cm) and samples at a depth of 20-25 cm were secured. Water was collected from the shallow tailings beaches around the vegetation stands. Sample Preparation The roots of the plants were washed free of tailings with distilled water. The resulting thick slurry of tailings was allowed to settle in the wash basin for five minutes and the supernatant water was decanted. The remaining saturated tailings were brought to 400 ml volume with distilled water in a beaker. The slurry was mixed with a magnetic stirrer for two 24-hour periods. After the first 24-hours, the slurry was allowed to settle for 20 minutes and the supernatant water was removed. The remaining sediments were suspended again in distilled water (400 ml total mixture) and leached for a second 24hour period. The tailings slurries from the different locations had a solid to liquid ratio, which ranged from 0.5 to 1.4 grams of dry tailings material per millilitre. The variations in the ratios are the result of differing fractions of coarse and fine tailings on the sites. In the final leachates, the solid to liquid ratios of the samples were 0.17 (±0.1) g/ml. The pH of the surface water and the leachate was determined with an I.L. Portomatic pH meter. All the tailings samples were homogenized in a mortar and brought to dryness at 75 to 85°C. Approximately 0.5 g of tailings were

Jan 1, 1981

Due Dilligance does not only lie in the statistics, but rather in the practical process aspects, which is the window to understanding the statistics.

Jan 1, 2005

By Frigyes Egerer, Pál Bárczy

Introduction Charles III, Holy Roman Emperor and King of the Austro-Hungarian Monarchy, established one of the first mining schools in Europe on June 22, 1735, in Selmecbanya (Schemnitz in German and Banska Stiavmica in Slovakian). The old city, built around the ancient gold mines, is shown in a contemporary engraving in Fig. 1. Charles' historical decree consisted of 13 chapters and included: the deed of foundation, the definition of grants, instruction for educational policy and curriculum, disciplinary rules, and the appointment of personnel. The first professor of the school, Samuel Mikoviny, gave his first lecture the same year, in December 1735. The 250th anniversary of the establishment of the School of Mines was celebrated Sept. 2-4, 1985, in Miskolc. It was attended by representatives of other Hungarian and of several foreign technical universities. A silver medallion (Fig. 2) was issued to commemorate the occasion and a series of lectures

Jan 6, 1986

By Michael G. Nelson

A computer simulation of the automatic control of continuous mining machines has been prepared, similar to previous simulations of longwall mining systems and coal thickness sensors. The simulation allows for programmed variation of the horizon detection method, the azimuthal guidance system, the type and configuration of the miner, varying hardness in the unmined face material (leading to varying error in the machine heading), and the miner's cutting cycle. It simulates control of the miner in three axes. Use of the simulator makes it possible to anticipate problems arising from the complex character of underground mining, which include unpredictable multi-pass coupling and unpredictable variation in geological conditions.

Jan 1, 1993

By E. Lefeuvre

Precambien in age, the phospate deposit under study which outcrops to the south forms a synclinal consisting of only one mineralized horizon. The synclinal strikes SSW/NNE, dips from 30 to 60 NNE and is about 5 km wide. The deposit has been explored by : - trenches and pits about 50 m apart in the outcropping area ; - core drill holes forming lines perpendicular to the outcrop. These lines are about 100 m apart. On each line, the drill holes are about 50 m apart The phosphate layer is clearly visible and has then been defined from the geologicia1 description of drill hole logs. The layer has been divided into two parts : a very rich, altered part at the hanging wall and a poor part displaying opposite physical characteristics, i.e. a fresh and hard rock. With open pit mining in view, the deposit has been divided into a 3-dimension network of blocks, each block measuring 50 X 50 X 100 m. The tonnages of rich and poor ore, PO25, MgO, Si02, CaO and R2O3 grades and the tonnages of barren waste at the hanging wall and the footwall have been calculated for each block. Given the dip of the phosphate layer ranging from 30 to 60° and its relative thickness (between 10 and 15 m), it was impossible or inappropriate to carry out direct calculation by 3-dimension kriging : too many neighboring blocks to obtain sufficient date and impossibility to draw up the morphology of the layer.

Jan 1, 1986

By Hans G. Machel

In low-temperature diagenetic environments (0 c T c 60 - 80°C), elemental sulfur forms by six common processes: via (1) inorganic oxidation of H2S by molecular oxygen, (2) inorganic oxidation of H2S by ferric iron compounds, (3) polysulfide disproportion- ation; and via the microbial metabolic actions of (4) colorless sulfur bacteria, (5) colored sulfur bacteria, and (6) some types of thiobacilli. Economically viable deposits of native sulfur usually are formed by only one process: inorganic oxidation of H2S by molecular oxygen. The amount of elemental sulfur is also limited by three common processes that destroy it in low- temperature diagenetic environments: bacterial reduction, bacterial oxidation, and inorganic oxidation. In high-temperature diagenetic environments (80 - 10o°C c T c 150 - 200°C), particularly in sour gas reservoirs, there are seven common processes forming elemental sulfur: (1) SOdH2S reaction, (2) H2S- hydrocarbon reaction, (3) polysulfide disproportionation, (4) catalytic H2S decomposition, (5) C02-H2S reaction, (6) chlorinification, and (7) exsolution of physically dissolved elemental sulfur. Natural processes that destroy or remove elemental sulfur in such settings are hydrocarbon sulfurization, polysulfide formation, and advection. In most sour gas reservoirs, relatively small quantities of elemental sulfur occur as a free phase (solid or liquid). However, elemental sulfur often precipitates in copious amounts from the reservoir fluids via exsolution and polysulfide disproportionation during production of sour gas reservoirs. By extension, elemental sulfur may precipitate from reservoir fluids in relatively large quantities upon uplift and erosion of deep sour gas reservoirs, or from advection of sulfurous fluids into shallower settings via faults or other high-permeability pathways.

Jan 1, 1992

By K. L. Bickel

Catalyzed diesel particulate filters (CDPFs) are used for reducing diesel particulate matter (DPM) concentrations in underground mines. The U.S. Bureau of Mines is conducting research on how catalyzed diesel particulate filters perform. Two CDPFs, with two different catalysts, are used on vehicles in two mines. The CDPFs are periodically removed from service and evaluated in the laboratory to determine their DPM collection efficiency, regeneration temperature, and effects on gaseous hydrocarbon (HC) and carbon monoxide (CO) emissions. This report summarizes the results of the laboratory evaluation, and the mines' experience with the CDPFs. Both CDPFs have operated over 2,000 h. One was evaluated after 1,646 h; its collection efficiency at one operating condition was 87.3 pct. After 2,135 h of operation it was removed from service because it was damaged. The second CDPF continues to be used after 2,328 h of service. The laboratory evaluation indicated its collection efficiency decreased after 1,584 h of use and its regeneration temperature increased from 405' C to 420' C. The mines' experiences with the two CDPFs varied. While being used on mine vehicles, both CDPFs became plugged with DPM on a number of occasions because of insufficient regeneration. CDPFs are capable of lowering DPM concentrations in diesel exhaust. They can also lower emissions of CO and HC, depending on the type of catalyst used. In this study, the performance of the CDPFs changed with use. Recommendations are given in this report for proper use of CDPFs in mines.

Jan 1, 1994

By David A. Collins

Red Dome Gold Mine classifies ore types within the pit by gold and copper grades. Copper grades are further distinguished by total copper and cyanide soluble copper levels. At present, Cyanide Soluble Copper samples take six hours of agitating before they can be assayed. This agitation time produces results which are a proportion of the total cyanide amenable copper. This relates to our plant's performance. Frequently the total copper result is needed before we can begin the cyanide soluble copper analysis. This caused a protracted tum around time. Mennetech Pty. Ltd has produced a catalyst proven by David Martin Menne to accelerate the cyanidation of gold. David said that the information at hand indicated that Leach WELL TM also accelerates the leaching of cyanidable copper substantially. Test work was undertaken on-site at Red Dome Gold Mine to investigate the potential of Leach WELL TM to reduce the leaching period and therefore, the overall tum around time required for cyanide soluble copper determinations.

Jan 1, 1995

By J. P. Savely

This paper describes the development of a trim blast design to improve the stability of a final pit wall. Description is given of the effects of rock characteristics, rock type, and pit plans on the blast design. Currently, six benches have been mined with trim blasting. A more stable final wall has resulted and the successful program has reduced mining costs over an extended time period. Low-cost blast monitoring and field observation were used to determine peak particle velocity and blast damage. Using criteria developed from this relationship, trim blasts were designed and evaluated.

Jan 1, 1987

By B. D. Hong

The Coal supply and demand review arrived too late to appear in the May issue. In 1991, the US coal industry produced 902 Mt (994 million st) of coal, according to preliminary data compiled by the Energy Information Administration (EIA) (Table 1). This amount was 32 Mt (35 million st) less than the record 933.5 Mt (1 billion st) produced in 1990. Most of the decrease in coal production occurred east of the Mississippi River. The substantial stock buildup that stimulated production in 1990 was not repeated in 1991. In addition, coal consumption by electric utilities and coke plants dropped by about 1.8 Mt and 4.5 Mt (2 million and 5 million st), respectively. Moreover, a higher level of coal exports failed to bolster the 1991 production. Production Surface mines accounted for about 59% of the 902 Mt (994 million st) of coal produced in 1991. Underground mines produced for 41%. More than one-fourth of the total coal output was from federal and Indian coal leases. Production of all ranks of coal, except subbituminous, declined in 1991. Bituminous coal is produced mainly in the East. It totaled an estimated 590 Mt (651 million st), a 6% decrease. Production of subbituminous coal, mostly in the West, rose 3% to a record 228 Mt (252 million st). Lignite output was from Texas and North Dakota. It decreased by 1 % to 79 Mt (87 million st). Anthracite production, all from eastern Pennsylvania, fell by 17% to 2.7 Mt (3 million st).

Jan 1, 1992

By S. R. Wardwell, M. J. Robison

INTRODUCTION The first phase of the Los Angeles Metro Rail Project consists of 7.1 km of subway and stations running through downtown Los Angeles from Union Station to MacArthur Park. Contract A146, Tunnels from 5th/HILL Station to 7th/Flower Station, involves tunnel construction through the heart of downtown Los Angeles. The alignment of the tunnels passes beneath twelve buildings, ranging from a 3-story below grade parking structure to numerous 12 to 14 story buildings. The soil along the alignment is alluvium, consisting of silts and clays, clean sands and gravels located above the ground water table. Because of unstable ground conditions, which led to excessive ground losses in several instances, and because of the potential for excessive building settlements, it became necessary to design and perform a program of chemical grouting. The chemical grouting program was designed such that tunneling could proceed on a one shift per day basis, with chemical grounting being performed on the remaining two shifts. Each day 21,000 to 31,000 liters of sodium silicate grout were injected through sleeveport grout pipes installed 18 m deep through the tunnel face. Tunnel advance rates of 12 m per day, 60 m per week, under full chemical grout cover were achieved. The chemical grouting program virtually eliminated both ground loss and settlement and permitted successful excavation of the tunnels. The purpose of this paper is to describe the chemical grouting program used to control ground losses and settlement on Los Angeles Metro Rail Contract A146. The paper begins with a description of

Jan 1, 1991

By Jeff Salai

The CP-3A contract is the final segment of the MWWST’s 28.3 km (17.6 mile)pressurized water supply tunnel for the metropolitan Boston area. The 1,288 m (4,226 feet) of U-shaped tunnel alignment passed through five rock zones employing four different primary support systems. The short tunnel began in Weston, Massachusetts, at Shaft 5A in soft, completely weathered argillite, passed through a fault zone, and ended in Dedham granite. An 18-foot diameter horseshoe starter tunnel was built to accommodate a 4.94 m (16-foot 2 1 / 2 inch) diameter Robbins TBM. Heavy water inflows were anticipated and subsequently encountered during probe drilling. Pre-excavation grouting was implemented at high pressures to displace high water inflows in the starter tunnel and consolidate the fault zone. This paper will provide an overview of the tunnel construction with an emphasis on ground control, grouting and survey control.

Jan 1, 2001

By C. B. Manula

This paper presents the results of an application of the PSU/USBM Open Pit Materials Handling Simulator (1) to a large stripping mine in the Southwestern Coalfield. The specific objective of the study was to demonstrate the application of the simulator to a complex multi-pit operation, with emphasis on simulating short term and long term production plans recommended for the mine by Theodore Barry Associates (2). Mine Description The mine is located in the Southwestern Coalfield, which encompasses parts of the states of Utah, Arizona and New Mexico. The lease area presently covers 121500 x 103 m2 (30000 acres) in a strip 37.007 x 103 m (23 miles) from north to south and up to 1609 m (1 mile) in width, located at an elevation above 1525 m (5000 ft). Three gently dipping coal seams are present, of which the upper seam (#3) averages 0-5.5 m (0-18 ft); the intermediate seam (#2) averages 0-1.2 m (0-4 ft); and the lower seam (#1) averages 0-2.4 m (0-8 ft). Overburden is mixed sand, gravel, calcareous clay, caliche and sandstone, with calcareous shale and sandstone burden between seams. Where the seams are covered by less than 6.1 m (20 ft) of overburden, they are generally weathered and oxidized and are not extracted. Stripping ratios have ranged from 3.5 to 6.0 with an average of 5.4. All future plans were based on a stripping ratio of 5.3.

Jan 1, 1979

Stope width was controlled around 3.5 feet for the narrow high grade vein system at Hollister Gold Mine. Without top access, backfill material was pumped from bottom access up into stope through a series of pipes after ore had been slushed out in each mining cycle. Backfill material was changed from one layer of low strength concrete plus one layer of high strength concrete cap to one layer of low strength cellular concrete plus one layer of high strength concrete cap. Backfill cost was cut down to 30 percent during this process. To reduce the steel concrete pipe set up and tear down time which was about 2 to 3 shifts in each mining cycle, steel pipes were left in man-way between mining cycles. Cellular concrete was firstly introduced into backfilling stopes at Hollister mine in Nevada Mining. This process including concrete pump selection, pumping pipe selection, material mixing, material mechanics and quality control were fully covered in this paper. CUT AND FILL MINING CYCLE The most flexible mining method, cut and fill (Figure 1), was used to mine the high grade, narrow steep vein system on this Northern Nevada property which is located eighty miles north east of Winnemucca. Without top access, stopes were accessed through two man-ways on each end, in which people travel ladders and material transport slides were set up (Figure 2). Each mining cycle was called a cut, from 6 to 8 feet high. Overhand sequence was used. Mining cycle started with drill and blast the middle portion of the stope, then expanded towards both ends. Broken ore was slushed into mill hole located at middle of the stope and then mucked out through bottom access into main haulage. Stope length was about 300 feet, which was restricted by slushing distance. Timber man-ways and mill hole would be raised and established for backfill and ready for the next mining cycle.

Jan 1, 2011

By Donald J. Olson, Thomas J. Lutzenberger, John Ramage

The North Shore Inline Interceptor, Phase IA tunnel is a 8,550-meter (28,000-ft) interceptor tunnel currently being constructed by Traylor Bros./Frontier-Kemper, Joint Venture, using a 9.8-meter (32-ft - 4-in) diameter Robbins tunnel boring machine (TBM). This tunnel is being driven in Silurian age dolomite approximately 91.5 m (300 ft) below ground surface. Significant water inflows were encountered when the tunnel drive penetrated the thinly-bedded Waubakee Formation. Flow increases approaching 0.75 cu m per min per 30 m of tunnel (200 gpm per 100 ft) have occurred in this formation. After driving approximately 2,285 m (7,500 ft) of tunnel (200 m full face in the Waubakee Formation), total flows exceeded 12 cu m per min (3,200 gpm) with about 5.7 cu m per min (1,500 gpm) entering the tun¬nel from the 200 m (600 ft) of exposed Waubakee Formation. Additional water control measures consisting of increased pumping capacity, increased water treatment capacity, surface grouting ahead of the tunnel, grouting ahead of the tunnel face and panning to improve the working environment for installation of tunnel support, were implemented to deal with the large water inflows. Currently (January 24, 1989), 3,400 m (11,000 ft) of tunnel has been driven, total water inflows are in the range of 11.7 cu m per min (3,100 gpm) with only 6.8 cu m per min (1,800 gpm) inflow from about 1,220 m (4,000 ft) of tunnel driven full face through the Waubakee Formation.

Jan 1, 1989