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By Stephen Pogorski, Colin R. Phillips, Atika Khan

An optimized counting scheme for a mixture of radon daughters and thoron daughters typical of Ontario mines, developed in a previous study, was experimentally examined under mine conditions. The counting scheme requires a 10 minute sample and 5 1/2 hours counting instead of a 60 minute sample and more than 40 hours counting as required by a previous method. Taking into account the uncertainties associated with concentrations, flowrate and radioactive decay, the precision of the counting scheme was analyzed and found to be on average better than the 40 hour method.

Jan 1, 1981

By H. Schultheis

The European Commission has launched in April 2002 the Thematic Network NESMI with-in the FP5 under Growth. NESMI stands for ?Network on European Sustainable Mining and Processing Industries? and will run until 2005. Within NESMI there are different ways to set up indicators and evaluate them throughout the NESMI Members. First analyses concerning the Demand and Competence within NESMI are developed. New focus points for analyses can be implemented into NESMI.

Jan 1, 2003

By V. N. Sukhov

This set was developed to use the mathematical methods and computers in the data analysis and accumulation of the large amount of geological data to solve the typical geological problems at the deposit exploring stage as well as for planning and managing in the open cast deposit mining. That is to say, the set of tasks ensures the geological data base at the stage of the deposit exploring and also its further use and. updating at the exploration stage. The following main elements may be ear-marked in the complex: (1) geological data base (GDB), (2) the data input system, (3) GDB maintainance, (4) "Geologist" geological application package (GAP). Let us consider now each element purpose. 1) GDB logical structure is determined through five hierarchical substructures with a common root segment.

Jan 1, 1986

By L. A. Obasi, C. Obasi

Investigations were carried out to determine the effects of changing metamorphic conditions on the formation of magnetite from the Itakpe iron ore deposit. Various magnetite grain sizes from ferruginous quartzites from different zones were determined and analyzed statistically. The results reflect the reaction of mineral individuals. In particular, the granulometry of magnetite changed with changes in the mineral formation conditions in the deposit and correlate with the mineralo-geochemical zones. This is interpreted as part of a readjustment of the mineral to the prevailing conditions. Because granulometry is a prebeneficiation parameter, this work will be, no doubt, of importance to the exploitation and beneficiation of the quartzites and rich ores.

Jan 1, 2000

By F. F. Aplan, C. E. Raleigh

Three run-of-mine bituminous coals, ranging in rank from low volatile to high volatile A, were floated at nominal -28 mesh (590 µm), -100 mesh (150 µm), and -400 mesh (37µm) feed particle sizes with selected pyrite depressants and a dispersant. For the three coals tested, pyrite and ash -forming minerals were usually rejected best at -100 mesh feed particle size. Depressant additions tended to improve flotation selectivity between pyrite and coal. When joint dispersant-depressant systems were used at -100 mesh (150 µm) feed particle sizes, the total sulfur content of the clean coal decreased further. For a comparable coal yield, less fuel oil collector was required for the flotation of -100 mesh coal than was required for the flotation of either -28 or -400 mesh feed.

Jan 1, 1992

By Jin Anshi, Wang Dai, Wang Zingshen

Both fire and gas problems in Fully Mechanized Sublevel Caving Coat Face are quite different from those in common fully mechanized coal face. Pouring N2into the goaf would be the best method to prevent fire in the goaf. In this paper how to determine parameters of pouring N2 is discussed. Comparing distributions of gases and temperature before pouring N2into the goaf with those after pouring N2 into the goaf, the effectivenesses of preventing the fire by pouring N2 into the goaf are testified. Some special gas problems appear in Fully Mechanized Sublevel Caving Coal Face, which are proved to be serious and interrelated with pouring N2 into goaf. Finally, a few methods adopted to deal with gas problems are described simply.

Jan 1, 1991

By R. R. Graham

Echo Bay Minerals operates a 650 tpd gravity float gold circuit at their Manhattan facility. The objective of this paper was to document the gold recovery by size distribution. This was accomplished by studying the material balance of each size fraction as it progressed through each gravity Circuit separator. It proved to be an impossible task. Figure I depicts the gravity circuit. The gravity cyclone underflow goes to two duplex Pan American type jigs. The jig overflow reports to the OK Flash Float Cell. The cyclone hutch product is reground and goes to a full size wilfly table. The rougher table concentrate reports to a pilot size Wi1fly cleaner table. The cleaner and rougher table tails circulate back to the jig feed. The gravity circuit was composite sampled for two to four hours per day for four days. Each sample was taken every 20 minutes. Each composite sample was filtered, dried, repu1ped, and screened through a wet/dry screening procedure. Each size fraction was checked for gold value by fire assay fussion technique. A material balance was run on each size fraction. The tables listing the results are enclosed.

Jan 1, 1989

By K. Kavouridis

The competitiveness of power production from lignite depends on the ability of coal industry to meet CO2 emissions limits, as specified by the Kyoto protocol. According to the National Allocation plan issued by EC, power plants should reduce CO2 emissions to a certain level, other-wise they will be obliged to cover the difference buying CO2 certificates through the Emission Trading Scheme. These charges will increase the operating cost of current units, as well as the investment cost for new units. To predict the increase in the marginal cost of electric power produced from the current plants of Western Macedonia, an emissions factor (CO2 t/MWh) for various units has been computed. Based on the algorithm used, this factor depends on the calorific value of the lignite, the efficiency of the power plant and the oxidation factor of the unit. For lignite plants with efficiency ~37%, CO2 emissions factor ranged from 1.19 t/MWh to 1.32 t/MWh. The increase in the marginal cost of kWh, as compared to the values of CO2 certificates, is higher in units fed with low quality lignite and worsens if the efficiency of the unit is low too. Also the presence of calcareous minerals in lignite increases the CO2 emissions and consequently the cost of power generation.

Jan 1, 2006

By Gary W. Hausler

Bergbau AG Niederrhein, a subsidiary of Ruhrkohle AG; Essen, Germany, is working coal mines on both sides of the Rhine River. The Prosper Mine in the City of Bottrop has a production of about 13,000 tons per day and has been in operation since the turn of the century. Mining has moved to the north as coal deposits south of Bottrop are exhausted. To facilitate ventilation, material handling, and transportation new shafts are being constructed in the northern coal fields. The Prosper 10 shaft is being sunk to develop the Prosper - Haniel coal field which has reserves of 280 million tons. In October, 1976 Bergbau AG Niederrhein awarded the shaft sinking contract for the 1047 m deep, 8 m inside diameter Prosper 10 shaft to Gewerkschaft Walter; Essen, Germany. Construction costs will amount to about 40 million DM with construction completion scheduled for 1981. GEOLOGY The shaft, located near the town of Kirchhellen north Bottrop, is being sunk in the Grafenwalder Horst between two faults on the south west flank of the Kirchheller Anticline. 270 m of upper cretaceus sands and marls overlay sandstone bedrock at the shaft location. The upper 124 m of this section contain highly saturated sands interspersed with marl. In order to successfully sink the shaft through this section, ground stabilization by freezing was used.

Jan 1, 1979

By S. McMurray

INTRODUCTION Towards the end of the 1890s the attention of a prospector by the name of Thomas Cullinan was drawn to the occurrence of alluvial diamonds east of Pretoria. Persistent prospecting work by Cullinan led to the loca¬tion of the source of these diamonds-a kimberlite pipe on the Elandsfontein farm. Cullinan eventually acquired the property in 1902, and during an extensive evaluation program outlined the largest diamondiferous pipe in South Africa on which the Premier mine was established. Worldwide attention was drawn to this mine in 1905 when the Cullinan diamond, which weighed 3106 carats and is by far the largest gem diamond ever discovered, was recovered. Premier mine operated with varying success until 1932 when the worldwide depression forced it to close. In 1945 the mine was reopened and since then has been a major producer of gem and industrial diamonds. During the early 1950s exploratory drilling revealed that below the 370-m level the pipe had been completely cut off by a younger gabbro sill. This sill has created major mining and metallurgical problems but the existence of vast reserves of high grade ore below it justi¬fied the establishment of virtually a new mine below the sill. Detailed planning of below-the-sill mining is nearing completion and the development of the first mining block (the L1 block) is well advanced. A full discus¬sion of this block follows. GEOLOGY OF PREMIER MINE Pipe Morphology and Country-Rock Geology As can be seen in Fig. 1, Premier mine is an elon¬gated oval shape with a long axis of 900 m on the surface and a short axis of 450 m. The pipe has a surface area of 32 hectares which decreases progres¬sively with increasing depth, so that 500 m below the surface the area is reduced to 22 hectares. The contact between the kimberlite and the surrounding country rock is sharp, with an average angle of dip of 1.48 rad (85°). From the surface down to approximately 350 m, the country rock is a felsite which grades downwards into a norite. In general both rock types are hard and massive, the felsite in particular being poorly jointed. This characteristic combined with other factors results in an extremely stable open pit, and near-vertical side¬walls are currently being maintained. The norite in general is competent but is heavily jointed in some areas resulting in very blocky ground which creates localized tunnel support problems. In the southeastern area below the sill a zone of highly altered and very unstable norite, which has also created localized sup¬port and development difficulties, lies adjacent to the pipe contact. Kimberlite Geology The Premier pipe is a complex multiple intrusion which contains in the region of 15 separate types of kimberlite, most of which are volumetrically insignifi¬cant. On a simplified basis three major types can be identified which correspond to three separate phases of intrusion. They are: Brown Kimberlite: This represents the first phase of intrusion and now occupies the eastern part of the pipe. In plan the brown kimberlite has a crescentric shape resulting from the intrusive relationship with the younger gray kimberlite which lies to the west. The brown kimberlite increases in relative area with increas¬ing depth which is a favorable factor for future mining operations since this kimberlite is the richest area of the mine, having an average grade of 70 carats/ 100 t. Below the sill this rock disintegrates extremely rapidly when exposed to air or water, to form a fine gravel-like material. This characteristic is due to a high montmorilonite clay content. As is common with clay minerals of this type, water can be absorbed into the crystal lattice with a resultant increase in volume which causes the physical decomposition of the rock. The presence of this decomposing kimberlite was a major factor governing the choice of a mining method, as well as development and tunnel support techniques. Gray Kimberlite: This is the most abundant type and represents the second major phase of intrusion. The diamond content is variable depending on the amount of waste rock dilution, and in general the rock is stable and non-decomposing. At the time of formation of Premier, the surface rock type was Waterberg quartzite, a rock formation which has subsequently been removed from this area by erosion. The explosive extrusion of the gray kimberlite caused extensive brecciation of this rock and while the kimberlite was in a mobile state huge masses of quartz¬ite, amounting to tens of millions of tons, slumped into the pipe. This quartzite became concentrated in the central area of the mine and has been a major mining problem throughout the history of Premier since, due to the irregular distribution of the quartzite, effective selective mining has never been possible. The quartzite is therefore extracted with the ore and has a significant dilution effect on the overall grade. Black Kimberlite: In the western part of the pipe a circular, pluglike body of kimberlite is intruded into the gray kimberlite. This plug consists mainly of a hard, dark-colored rock known as black kimberlite, but also contains a number of minor kimberlite variet.ies as well as nonkimberlite carbonate dikes. The Gabbro Sill The sill averages 75 m in thickness and within the confines of the pipe consists of 52 million t of rock. On average the major dip is 0.34 rad (20°) to the northeast which is across the short axis of the pipe, the range of dip being from near horizontal up to 0.52 rad (30°). Due to the pronounced dip, the sill is encount¬ered over a vertical distance of 175 m with the highest upper contacts on the 355-m horizon and the lowest bottom contact on the 530-m horizon.

Jan 1, 1982

By H. Wagner

INTRODUCTION One of the most important tasks facing the opera¬tors of deep-level gold mines is supporting the roof strata in the immediate vicinity of the stope faces. Some of the deeper gold mines operate at depths ex¬ceeding 3.0 km; the extreme mining depths and the tabular nature of the gold-bearing reefs result in high stress concentrations at the abutments of the stope faces. These, in turn, lead to extensive fracturing of even the hardest quartzites. Although most of these fractures develop in a stable manner, there have been many instances where rock failure occurred suddenly, resulting in a considerable amount of damage to the underground workings. Such violent rock failures commonly are known as "rock bursts," and they are typical of many hard-rock mines that are operating at great depths. To safeguard mining personnel and equipment against the effects of stable and unstable rock failures, adequate support must be provided to the fractured roof as quickly and as effec¬tively as possible. Traditionally, timber props and timber packs have been utilized to provide roof support in gold mine stopes up to 1.5 m wide. Both methods of support suffer from the fundamental disadvantage of generating sup¬port only in reaction to movement of the strata. Since unrestrained strata movement often results in a break¬down of the frictional forces that preserve the stability of the intensely fractured roof strata, it is essential that stope supports have rapid load-bearing characteristics. An equally important requirement is that the roof sup¬ports be able to accommodate displacements of the rock mass. A third requirement of roof supports in gold mine stopes is that they must be able to accommodate the high rates of stope convergence typical of rock bursts. The hydraulic prop is the obvious choice for sup¬porting the stope faces in deep-level hard-rock mines. Since hydraulic props can be installed with a positive setting load, they can provide roof support that is independent of strata movement. By equipping the hydraulic props with yield valves, they meet the require¬ment of being able to accommodate rock-mass displace¬ments. However, the requirement of being able to accommodate high rates of stope convergence without overloading or damaging the rock strata was rather difficult to meet. The first introduction of hydraulic roof supports into gold mine stopes showed that conventional hy¬draulic props, such as those used in coal mines, were completely unsuitable for a number of reasons, in¬cluding: 1) Since drilling and blasting are used to break the ore in gold mine stopes, supports installed close enough to be effective are exposed to the vigor of the blast. Conventional coal-mining props suffered extensive blast damage, despite being protected by a blast barricade. 2) The environmental conditions in gold mine stopes favor corrosion. The high quartz content and the abrasiveness of the ore further aggravate the problem by damaging the protective coatings on the hydraulic surfaces. Simply changing the design of gold mine props to have the piston project from the bottom rather than from the top resulted in dramatic improvements in the service life of the props. The improvement was attributable to external water in the stope washing the fine quartz grit away from the hydraulic surfaces, rather than concentrating the particles between the cylinder and piston of the prop. 3) The most obvious deficiency of conventional coal mining props was their inability to yield rapidly enough to accommodate the high convergence rates of rock bursts without overloading the supports. As a result of the first disappointing results with hydraulic supports in narrow hard-rock stopes mined by drilling and blasting, a comprehensive research and development program was undertaken. The hydraulic props resulting from several years of research and exten¬sive field trials are described herein. To date (1978), more than 130,000 of these props have been manufac¬tured and used successfully in a number of mines and under a variety of geological conditions. Although many of the detailed specifications are directed spe¬cifically to gold mining conditions, there are a number of points that are of general importance in the design of underground support systems for tabular excavations in hard-rock mines of any type. SPECIFICATIONS FOR RAPID-YIELDING PROPS Yield Requirements The yield requirements for hydraulic props in deep¬level hard-rock gold mines are summarized in the fol¬lowing paragraphs. Slow Yield Load: The design of the yield valve must allow the yield load to be preset to between 300 and 400 kN. This flexibility is necessary to account for the differences in the strength properties of the roof and floor strata. Drop in Load Under Slow Yield: The drop in sup¬port load during a slow yield should be as small as possible, but it should not exceed 5% of the slow-yield load. Rapid-Yield Load: The load exerted by the prop during periods of rapid stope convergence should be as close as possible to the slow-yield load. The rapid-yield load should not exceed the slow-yield load by more than 15% at yield rates of 1.0 m/s. The limitation of the upper valve of the yield load is important; it deter¬mines the hydraulic performance of the yield valve. The hydraulic props that satisfy this criterion have performed well, even under severe rock bursts. Load Changes Under Rapid Load: On some of the earlier rapid-yielding props, conditions of dynamic in¬stability were observed; these conditions resulted in a complete breakdown of support resistance. It is essen¬tial that the force oscillations during rapid yielding be as small as possible, not exceeding -*5% of the average rapid-yield load at a yield rate of 1.0 m/s. Longitudinal Resilience: During rock bursts, vibra¬tions of the roof and floor sometimes are observed. To

Jan 1, 1982

By J. F. Mack, P. R. Michael, J. M. Shapaka, S. J. Self, M. W. Richmond

"On October 11, 2000, an estimated 306 million gallons of water and coal waste slurry drained from an impoundment in Martin County, eastern Kentucky into an adjacent underground mine. Approximately 230 million gallons of the water and slurry discharged from two underground mine portals and affected over 75 miles of streams in Kentucky and West Virginia. In response to this and several other similar events, the U.S. Office of Surface Mining Reclamation and Enforcement (OSMRE) commissioned an oversight study that evaluates how well state and federal regulatory programs established under the Surface Mining Control and Enforcement Act are ensuring that impoundment operators are minimizing the potential for impoundment-basin breakthroughs. The study focuses on slurry impoundments constructed in the hollows of Appalachia where numerous coal seams and steep topography combine to result in a large number of mined seams intersecting and underlying impounding facilities. Factors under consideration include: (1) full accounting of all mineable coal seams intersecting and underlying the impoundment; (2) identification and accurate location of underground mines close enough to the impoundment to potentially affect its stability; (3) assessment of the stability of coal barriers between the impoundment basin and adjacent mines, and stability of roof rock and pillars in mines subjacent to the structure; (4) determination of the flowability of the impounded slurry when expanding or undermining the facility are being considered; and (5) measures undertaken to reduce breakthrough potential when necessary. The oversight study began in 2012 after program-evaluation criteria were defined in an OSMRE technical position paper. The study is scheduled to be completed in 2017. BACKGROUND The Surface Mining Control and Reclamation Act (SMCRA) authorizes the disposal of coal mine waste by constructing a coarse coal mine waste embankment across a valley and pumping the fine coal mine waste (slurry) into the resulting basin (Figure 1). A concern shared by many engineers, geologists, and mine inspectors familiar with coal mine waste slurry impoundments is related to the common occurrence of underground mine workings adjacent to or beneath the impoundment: the potential for slurry “breakthroughs” into mine works and subsequent breakouts into the surface waterways (Figure 2). On October 11, 2000 a combination of coal refuse slurry and water from the Big Branch impoundment in Martin County, Kentucky broke through into an underground mine and subsequently discharged into the receiving streams. The location of the point of breakthrough, pathways of slurry/water flow within the mine workings, and two points where the liquid exited the mine are shown in Figure 3. An estimated 306 million gallons of water and coal refuse slurry drained from the impoundment into the adjacent underground mine. Approximately 230 million gallons subsequently discharged from the underground mine at two portals."

Jan 1, 2015

By R. H. King, D. R. Schricker, R. E. Cameron

Mine managers rely on their experience, section foreman's daily reports and occasional time studies to make important production, maintenance and forecasting decisions. This information is often subjective and imprecise, consumes expensive engineers time, is often biased and is restricted to short time periods. Managers need better information. Furthermore, if a recently purchased machine is being evaluated, a large increase (>10%) in productivity may be obvious, but small increases are usually masked by other variables. In today's highly competitive markets, cumulative small changes are important, especially since managers have already exhausted most available, obvious avenues of improvement. A technique to separate these charges from effects of other variables and to quantify the target effect is necessary. CSM researchers are addressing this need by developing an intelligent decision support system (IDSS). It provides key production and maintenance information to support management decisions by recognizing patterns in mining machine sensor data and representing knowledge about mine management, operations, layout and equipment. A continuous miner section at Western Fuels-Utah Inc.'s Deserado Mine in Rangely, CO is the test facility for the research. Power transducers located in the sections load center provide kilowatt versus time data from continuous section equipment. The IDSS pattern recognition algorithms classify the kilowatt versus time patterns into individual machine operating modes. It produces an output similar to a typical industrial engineering time study but continuously, automatically and without an engineer in the section. It is relatively easy for a time-study engineer to identify the stop and end points of a cycle element and classify it by looking at the machine and glancing at a stop watch. The problem is that monitoring system hardware and software do not have these unique capabilities. Consequently, IDSS algorithms must automatically define the begin and end points of machine tasks and identify the task by recognizing patterns in kilowatt versus time data. However, many mines have file drawers of time-study type data that are not very useful. So, IDSS contains the knowledge representation subsystem (IDSS/KRS) to produce short reports containing only key information. This is difficult because mine environments continually change and information requirements differ for each management level. Consequently, IDSS/KRS must learn and adapt to new conditions in a dynamic mining environment. Western Fuels-Utah owns and operates the Deserado Mine, located about 19 km (12 miles) northeast of Rangely, CO (Upadhyay, 1988; King and Eros, 1988). One longwall and two continuous miner sections produce about 900 kt (1 million st) of coal annually for the Deseret Generation and Transmission Cooperative's 4000 MW power plant located near Bonanza, UT. Three mine monitoring systems control the underground mine, the preparation plant and the unit train load out. Pattern recognition IDSS/PRS produces a listing of begin and end times for each machine operating event as shown is Table 1.

Jan 1, 1991

By Jim Oehler

The US minerals industry, as well we know, has been hit hard by the recent recession and is faced with what is projected to be a slow recovery. Even so, there was an underlying feeling of optimism at American Mining Congress' International Mining Show in Las Vegas, NV, Oct 11-14. More than 25,000 people attended during a particularly rough financial year, dispelling fears that attendance would dip sharply below 1978's 30,000 plus tally. The equipment expositionheld every four years-was the largest show of its kind this year in the Western Hemisphere in terms of exhibit space, providing an extensive view of mining-related equipment. In addition to the nearly 700 exhibits occupying more than 445 hm2 (11 acres), 20 technical sessions were held covering topics ranging from minerals availability and the state of the mining industries to communications, computers, and accounting. Government officials were well represented, with six US senators and five US representatives participating in various convention functions. Sen. Paul Laxalt (R-NV) was the keynote speaker at the welcoming luncheon. Other senators involved in the show included Howard W. Cannon (D-NV), Pete V. Domenici (R-NM), James A. McClure (R-ID), Frank H. Murkowski (R-AK), and Steven D. Symms (R-ID). Representatives on the program were Joseph M. Gaydos (D-PA), Guy Vander Jagt (R-MI), Don Young (R-AK), Jim Santini (D-NV), and Ceil Heftel (D-HI). There were also many international vistors at the show, including exhibitors from Australia, Austria, Canada, England, Finland, Sweden, South Africa, and West Germany. Equipment Highlights Equipment and services featured at the exposition ranged from aerial mapping to wire rope. Aisle space alone between exhibits exceeded 8 km (5 miles). And the companies represented ranged from small businesses to large corporations. Some of the company products described here are representative of the many exhibitors in attendance. Atlas Copco's extensive display of rock drilling and tunnel and shaft boring equipment featured its Boomer H-201, the newest addition to its line of underground drifting rigs; the ROC 306, a new track-mounted, high pressure down-the-hole drill rig; a new hydraulic surface crawler drill, the ROC 712H, designed for small and medium-size holes in quarries and small open-pit mines; a computerized analysis system, CARE (see page 1570), designed to project productivity and operating costs for various mining operations; several electro-hydraulic loaders, including the new rail bound LM E60 and LM E80, and two Hagg-loaders-the 7HR and 10HR-that use two digging arms to rake and guide muck; a mechanized blasthole charging system; rock bolting equipment and its new Swellex roof bolting method; Sandvik Coromant drill steel and bits; and its new Jarva R200 raise boring machine.

Jan 11, 1982

By I. K. Choi

In a recent paper by Smith and Wen(1) the adsorption of sodium dodecylsulfate and trimethyldodecylammonium chloride on kaolinite was studied. As would be expected considering the general negative character of kaolinite, except perhaps at quite' acid pH values (2,3,4), it was found that trimethyldodecylammonium chloride adsorbed strongly on kaolinite at basic pH values. Furthermore, it was found that sodium dodecylsulfate adsorbed strongly on kaolinite at acid pH values, particularly at initial 1H values below pH 3 (final pH values near pH 4). It is interesting to note that adsorption of the anionic surfactant was strong under conditions where the kaolinite was apparently slightly negatively charged (2). The present paper reports on an extension of the work by Smith and Wen. In this paper the effects of the presence of dodecyl alcohol and sodium hexametaphosphate on adsorption of sodium dodecylsulfate and trimethyldodecylammonium chloride on kaolinite are considered.

Jan 1, 1967

By K. Whiting

In October 1993, the first two of a series of full scale passive biological treatment cells were constructed to treat mine drainage from the Burleigh Mine (located approximately 48 km west of Denver, Colorado in the Clear Creek/Central City Superfund Site). The two cells are each 19-21 m long, 10.7 m wide and 1.2 m deep. Each cell contains approximately 216 cubic meters of substrate (composted manure) and is designed to treat 60 L/min of mine flow. One cell is designed as an upflow reactor and the other cell is designed as a downflow reactor. The differences in flow hydraulics and treatment efficiencies will be tested in these two cell for approximately nine months before the remaining cells are built to treat all the discharge. The substrate material was selected based on a series of bench scale tests. During the tests, the amount of metals removed was determined. The design flow rate was based on a sulfide generation rate of 400 nanomoles of sulfide generated per cubic centimeter of substrate per day. Metal removal results for the first three months of operation of the two cells will be presented and compared to results of bench and pilot studies at other sites.

Jan 1, 1994

By Michael Fehler, Hidenshi Kaieda, Leigh House

Hydraulic fracturing has gained in popularity in recent years as a way to determine the orientations and magnitudes of tectonic stresses. By augmenting conventional hydraulic fracturing measurements with detection and mapping of the microearthquakes induced by fracturing, we can supplement and independent1 y confirm information obtained from conventional analysis. Important information obtained from seismic monitoring includes: the state of stress of the rock, orientation and spacing of the major joint sets, and measurements of rock elastic parameters at locations distant from the wellbore. While conventional we1l 1ogging operations can provide information about several of these parameters, the zone of interrogation is usually limited to the immediate proximity of the borehole. The seismic waveforms of the microearthquakes contain a wealth of information about the rock in regions that are otherwise inaccessible for study. By reliably locating the hypocenters of many microearthquakes, we have inferred the joint patterns in the rock. We observed that microearthquake locations do not define a simple, thin, planar distribution, that the fault plane solutions are consistent with shear slippage, and that spectral analysis indicates that the source dimensions and slip along the faults are small. Hence we be1 ieve that the microearthquakes result from slip along preexisting joints, and not from tensile extension at the tip of the fracture. Orientations of the principal stresses can be estimated by using fault plane solutions of the larger microearthquakes. By using a joint earthquake location scheme, and/or calibrations with downhole detonators, rock velocities and heterogeneities thereof can be investigated in rock volumes that are far enough from the borehole to be representative of intrinsic rock properties.

Jan 1, 1986

By K. B. C. Brady, B. E. Scheetz, J. E. Cuddeback, T. W. Bergstresser, W. A. Telliard, S. C. Parsons, R. J. Hornberger

Two weathering procedures are being developed by the Acid Drainage Technology Initiative in cooperation with federal agencies(OSM and EPA). Objects are to: (1) standardize existing humidity cell and leaching column procedures; (2) improve test methods by maintaining a carbon dioxide enriched environment needed to optimize carbonate mineral dissolution, and quantifying particle size variables to evaluate reaction kinetics, and (3) provide flexibility in test method implementation consistent with EPA guidelines for Performance-Based Measurement Systems. Preliminary results indicate the CO2enriched leaching environments produce alkalinities that are similar to field observations and greater than the tests run under atmospheric PCO2. Leaching columns produce higher alkalinity than humidity cells.

Jan 1, 2003

By R. B. Krog

Airflow patterns in longwall bleeder systems are affected by gob permeability and porosity and the condition of the gateroad entries after mining. As longwall panels are mined, areas of the caved zone and portions of the remaining gateroad entries become inaccessible, making direct ventilation measurement impossible. Mine operators are left to estimate direction and amount of airflow in the gob and the surrounding gateroads. The most readily available locations are where air enters the worked-out area, approved measurement point locations (MPLs) and/or evaluation points such as bleeder evaluation points, intake evaluation points, and the outer bleeder entry. Ventilation models (network or Computational Fluid Dynamics, CFD) can model the gob and the surrounding bleeder system but calibrating these models is difficult without field data. NIOSH installed a tube sampling system ahead of a longwall face and, as the face retreated, released a slug of tracer gas (SF6) to determine transportation times and flow rates in the inaccessible gateroads. The data was used to calibrate both network and CFD ventilation models for a longwall panel. Both model types were used to predict the airflow patterns in the gob and to determine the airflow rate from the longwall tailgate corner back to the bleeders.

Jan 1, 2011

By J. Engesser

A research project was undertaken to determine if coal addition tofluxed-magnetite pre-fired pellets would reduce nitrogen oxide (NOx)emissions during pellet induration. Iron oxidation rate tests, coal oxidation rate tests and calcination rate tests were conducted in a laboratory furnace to determine the amount of coal necessary to provide adequate heat for limestone calcination without producing detrimental effects to magnetite oxidation. Pilot scale pot-grate tests were then performed with and without coal in the pre-fired pellets to verify pellet quality parameters and nitrogen oxide emissions. The off-gas during each pot-grate test was analyzed for NOx, CO2, CO, SOx, and O2. There was approximately a ten percent reduction in NOx emissions with coal addition during the normal pot-grate firing cycle. There was a 75% reduction in NOx emissions during firing cycles that used 25% to 75% excess air during firing rather than the normal 300% excess air.

Jan 1, 2003