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By T. Y. Yan

O2, NaCl03, H202, NaOCI and Ce+4 have been examined as oxidants in an alkaline carbonate system for uranium in situ leaching. The South Texas and New Mexico ores contained up to 2.6 meq/g of reducing compounds that consume oxidants in the leaching operation. Leaching rates did not increase linearly with the oxidation potential of the oxidants. A mild oxidant can be used to maximize uranium oxidation selectivity and oxidant efficiency. 02 can oxidize most of the uranium mineral with an adequate rate and high oxidant efficiency.

Jan 1, 1993

By D. Tao

The reverse cationic flotation is the most widely utilized flotation method for concentrating iron ore. In such a process quartz is often floated with ether amines (R?O(CH2)3?NH2) partially neutralized with acetic acid. However, it is well known that the cationic flotation of hematite does not work well if a certain amount of slime is present in ore slurry. To improve flotation performance, starch is often added to the flotation feed to selectively flocculate and depress the iron minerals and sodium silicate is used to disperse the silica gangue. The present study was conducted to improve the flotation efficiency of iron ore flotation by use of a new type of depressant for clay and iron minerals, i.e., GP clay binder. It is postulated that the GP clay binder minimizes slime adsorption on quartz and hematite particles by agglomerating clay particles to reduce their surface area. It is further theorized that the clay binder helps depress hematite flotation due to its unique chelating ability to iron ion. The results from this study clearly show that the GP clay binder is an effective depressant in iron ore flotation. When it was used as a substitute for corn starch, it increased concentrate grade from 60.78% to 67.67% and increased iron recovery from 71.96% to 72.90%. The combined use of GP clay binder with corn starch produced even better flotation performance, increasing concentrate grade to 67.08% and iron recovery to 78.15% simultaneously. Key words: Clay Binder; Depressant; Iron Ore; Flotation.

Jan 1, 2009

By Justyna Swolkień, Nikodem Szlązak, Dariusz Obracaj

Polish coal mines are characterized by severe geological conditions and the co-occurrence of natural hazards. These hazards are associated with methane and dust emissions during mining processes which can provide methane, dust or methane-dust explosions. Besides, there are such natural hazards as gas and rock outbursts, rock bursts, spontaneous combustion of coal and environmental heat load. The mutual influences of various geological conditions, as well as the natural hazards, intensify the increase of the methane explosion hazard. The occurrence of methane explosions in the Polish mining industry between 2013 and 2108 are analyzed. Fatal accidents are compared with similar events in the global mining industry. The methods of preventing the explosion hazard as well as methane emission control in the Polish coal mining sector are presented. Methane hazard prevention should be adapted to levels and other natural hazards, especially those related to ventilation of underground mine. The methods of preventing methane emissions into mine excavations are often contrary to the means of preventing other ventilation-related hazards in the underground mine, e.g. the prevention of spontaneous combustion of coal in the goaf of longwall panel. Therefore, the design of longwall panels should be preceded by a detailed analysis of the results of both methane emission prediction and forecast of other ventilation-related hazards.

To build an effective entry roof beam, a roof bolting or supporting system must be of the ability of providing a clamping force to the bolted roof strata. Currently, two roof bolting systems are widely used to serve the purpose of providing such a clamping force. They are resin and mechanical anchorage bolting systems. Both numerical modeling and double shear testing results indicate that a greater pre-installed load or pretension of a roof bolting system is very critical for effective beam building. Although it?s generally accepted that a mechanical anchorage bolting system can provide a greater pretension than a resin anchorage bolting system does (Su and Poland, 2007 and Peng, 2008), no published field data are available to support such a conclusion. This paper analyzes the results of our field tests on both roof bolting systems, explains why one roof bolting system is better than the other under a typical roof condition, and finally discusses the effective means for coal mine operators to determine the best roof bolting system for their roof conditions.

Jan 1, 2011

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

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

Jan 12, 1986

By D. R. Wilburn

This summary of international mineral exploration activities for the year 2005 draws upon available information from literature, industry and U.S. Geological Survey (USGS) specialists. It provides data on exploration budgets by global region and mineral commodity and identifies significant mineral discoveries and exploration target areas. It also discusses government programs affecting the mineral exploration industry and presents analysis of the mineral industry based on these data. Two types of information are reported and analyzed in this annual review of international exploration for the year 2005. They include budgetary statistics provided by Metals Economics Group (MEG) and information on regional and site-specific exploration activities in 2005 compiled by the USGS.

Jan 1, 2006

By K. Norrena

Grade control in open pit mining requires the specification of dig limits that account for mineral grades, economic costs, and selectivity of mining equipment. Visual identification of ore rock types is ideal, but this is not always possible in lower grade disseminated deposits where ore/waste contacts are not visually discernible. In this case, conventional grade control practice consists of a two-step procedure (1) create a map of mineral grades at some selective mining unit scale, and (2) determine practical ore-waste boundaries or dig limits on the basis of the gridded block grades or assay information. This procedure is laborious, depends on a subjective assessment of where the boundary should be, and may be economically sub-optimal. We pose the determination of dig limits as an optimization problem and solve that problem with the technique of simulated annealing. Simulated annealing has the unique advantage of being able to combine multiple non-linear constraints into a single objective function. We use maximum profitability and the ability of the equipment to mine the proposed dig limits as constraints in the determination of optimal dig limits. We require a map of expected profit for eah block. Geostatistical techniques are recommended, for mapping expected profit but not necessary. Geostatistics will provide a quantification of the uncertainty in the grades within rock types using all available blasthole samples and exploration drilling. Some variant of L-optimal estimation or kriging can be used to determine the block-by-block classification that is economically optimum. We also need the expected profit for each block. It is unrealistic to assume free selection, that is, each block cannot be extracted independently of its neighbors. The optimal balance of ''accepting dilution'' and ''wasting ore'' is achieved to maximize profit subject to equipment constraints.

Jan 1, 2001

By Brianna M. Eiter, Jennica L. Bellanca

Mineworkers face a challenging and dynamic work environment every workday. To maintain a safe workplace, mineworkers must be able to recognize worksite hazards while they perform their jobs. Though hazard recognition is a critical skill, recent research from the National Institute for Occupational Safety and Health (NIOSH) indicates that mineworkers fail to identify a significant number of hazards. To further the understanding of mineworkers’ hazard recognition ability and to begin to address hazard recognition performance, NIOSH researchers analyzed data collected during a laboratory research study to determine the effect of individual mineworker factors including risk attitude, work experience, and safety training on hazard recognition accuracy. The results of this study show that mineworker risk attitude and safety-specific work experience affect hazard recognition performance while hazard-specific safety training does not. These results suggest that some of these individual factors can be overcome through experience and training. Potential strategies that can be used to address these factors are also discussed.

Aug 31, 2020

By F. B. Henderson

1.0 PROLOG Since the "end" of the Cold War, revolutionary new research and commercial remote sensing capabilities are being developed that will dramatically change the way we map the earth and address its environmental problems. The technology challenge is not the growth of new environmental remote sensing technologies themselves. It is, rather, how these new innovative observing systems can be accepted and used to the benefit of both the regulatory and the mining communities in the light of present concerns about unknown risks of discovery and liability. The basis for this paper is an unpublished technology primer prepared for the industry-government Workshop on Advanced Measurement and Site Characterization for Mining Impacts held at Colorado School of Mines, Golden, Colorado, July 14-17, 1997

Jan 1, 1999

By A. Nieto

A dump edge and vehicle proximity warning system is being developed at the Colorado School of Mines using GPS and wireless local area networks to improve the safety of off-highway trucks in open pit mines. After two years of research, software development, and testing in the laboratory, field tests are currently being carried out at operating limestone quarries and open pit mining operations to evaluate the effectiveness of the system under rugged operating conditions. This paper describes the progress made on this NIOSH-funded project and discuss, tests carried out recently, at the Morenci open pit copper mine in Arizona.

Jan 1, 2002

By Richard L. Bullock

Believe it or not, mine evaluation and feasibility studies are as old as the industry itself. In the first recorded writing on mining by Agricola (1556), he gave several clues as what to look for in evaluating a mine, to determine if it was economically feasible. Below is the guidance from Agricola, with my interpretation of what he would include today. Now a miner, before he begins to mine the orebodies must consider seven things, namely: ? The situation (the geography/topography). ? The conditions (the geologic/mineralogic and mining conditions). ? The water (availability, hydrology and ownership). ? The roads (plus the railroads, power lines, all other infrastructure). ? The climate (the weather-rainfall/environment). ? The right of ownership (the land and legal entitlement). ? The neighbors (the social environmental concerns of the area). It is amazing how much of his guidance has not changed in 455 years. Yet Agricola gives no guidance on determining the correct cost, or determining the correct commodity pricing. But he does give a few words of caution on the related ethics involved on buying mineral properties: ?A prudent owner, before he buys shares, ought to go to the mine and carefully examine the nature of the vein, for it is very important that he should be on his guard lest fraudulent sellers of shares should deceive him.? Three hundred years later (1866), to further amplify this point, in the most famous thing that Mark Twain never said, as an introduction to a lecturer to an audience of gold-miners at Red Dog, CA, the famed author/humorist allegedly described a mine as ?a hole in the ground owned by a liar? (Farrar & Rinehart, 1935). So, one might conclude that between these two dates, the mining industry did little to improve its reputation. Unfortunately, since Mark Twain allegedly made those remarks, we have not done near enough to gain accountability for the mineral industry.

Jan 1, 2011

By C. T. Agouridis

Head-of-hollow fills or valley fills have gained national attention due to increasing environmental concerns, particularly with regards to headwater stream loss. Researchers at the University of Kentucky in conjunction with outside scientists and consultants in the fields of stream restoration, wetland restoration, and mined land reclamation have developed new design methodologies for creating a headwater stream system for a head-of-hollow fill in eastern Kentucky. The design was largely built on the Forestry Reclamation Approach (FRA), which encourages a non-compacted spoil medium to promote tree growth, in an effort to address concerns related to water quantity and quality as well as habitat development. The major components of the design included 1) modifications to the crown geometry, 2) compaction of the crown to control infiltration, 3) utilization of natural channel design techniques, 4) use of the FRA to promote tree growth, 5) creation of ephemeral channels and vernal ponds, and 6) implementation of a novel bioreactor-wetland treatment system to improve water quality.

Jan 1, 2008

By Ed Cope

The newly discovered South Arturo deposit is a Carlin-type, breccia-hosted, oxide gold deposit located on the Dee Mine property at the north end of the Carlin Trend in Nevada. The Dee Gold Mine was discovered in 1981 by Dee Delamar, a Cordex geologist. The property was put into production by Rayrock Mines in 1984. During its 16-year mine life, the Dee Mine produced about 18.6 t (600,000 oz) of gold from oxide ore at an average production rate of about 1.24 to 1.55 t/a (40,000 to 50,000 oz/year). In October 1997, Bar-rick Gold Exploration acquired the right to earn a 60-percent interest in the Dee property from Dee Gold Mines through an exploration farm-out agreement. Barrick Gold Exploration began exploration activities on the property in 1998 and vested in its 60 percent operating interest in 2004. However, the discovery hole at South Arturo was not drilled until July 2005 from atop the historic mine dumps located 300 m (985 ft) south of the Dee openpit. The discovery hole (BD-72) encountered a cumulative 75.6 m (248 ft) of oxide mineralization grading 1.9 g/t (0.055 oz/st) gold beneath 260 m (855 ft) of dump material and post mineral Tertiary Carlin Formation volcanic and sedimentary rocks. Oxidation within the deposit extends to depths greater than 610 m (2,000 ft). The deposit has passed the prefeasibility test and is presently in the feasibility stage. Exploration is on-going to enhance or enlarge the known deposit. The current resource for the deposit stands at 40.4 t (1.3 million oz) of gold. The South Arturo discovery is an example of an historic mining property yielding a new discovery through the diligent application of area selection criteria of the model type and persevering when early results were generally negative. Barrick?s original exploration pro-grams on the property were focused on the discovery of deep, high-grade sulfide deposits that would likely be mined by underground methods. As it turns out, South Arturo is mostly oxidized and will be amenable to mining by openpit methods

Jan 1, 2008

By Shawn Collins

BACKGROUND The history of the Pine Point Mine is a rich one. This property once boasted one of the largest and most profitable lead-zinc mines in Canada's long and storied mining history. The Pine Point mine used open pit technology to extract the ore, but they faced several challenges. In addition to logistics problems with haul distances and metal prices, the mine faced the challenge of a large in flow of water from the Great Slave Lake. Eventually, the Pine Point Mine was closed leaving behind a large deposit of lead-zinc ore. Following the closing of the Pint Point Mine, Tamerlane Ventures Inc. acquired this property which includes approximately 70 million tonnes of lead-zinc reserves. The initial phase of this project will concentrate on the R-190 deposit as a test site to prove their new approach for extracting the ore. The R-190 reserve is believed to have approximately 1,014,000 million tonnes of high grade ore varying from 5% to 52% combined lead-zinc. Once the pilot project proves the approach is valid, the process will be repeated in approximately 33 other ore bodies on the property. This paper is written to discuss the design process to determine the most effective means of placing production and service equipment in the same shaft. Currently the design work and the mining scheme is complete, but lead-zinc pricing has prevented Tamerlane Ventures Inc. from moving forward with the project.

By M. M. Rajkovich

Just prior to the close of the century, the Federal Mine Safety and Health Review Commission issued its long-awaited decision in Berwind Natural Resources, et al v. Secretary of Labor, to usher in the new millennium. The case arose as a result of a mine explosion in Eastern Kentucky resulting in MSHA’s issuance of 225 citations, not only to the contractor, but to four affiliated entities of Berwind Corporation. While an Administrative Law Judge dismissed those citations as to three of the Berwind entities, both MSHA and the remaining Berwind entity appealed the case to the Commission. Although the Commission effectively left intact the Administrative Law Judge’s decision, it announced a new theory of liability that would apparently be applied in the future, namely the Unitary Operator Theory. This presentation will provide an in-depth look into the Berwind decision and explore the means under which MSHA might impose liability for contractor violations upon other entities in the future.

Jan 1, 2001

By Y. P. Chugh

During late 2010, a scientifically founded water spray system was developed for continuous miners (CM) (Patent Pending) that reduces respirable dust exposure for miner operator by about 30%. It includes engineered sprays around the cutting drum, outer bit-ring sprays, chassis sprays and side cutter boom sprays. About 3.6 ? 4.5m (12-15 ft) behind these sprays is additional sprays that spatially cover the area between the CM chassis and the mine roof and ribs. The spray system uses approximately the same volume of water as currently used on CMs with similar operating pressures. An Illinois mine decided to implement these concepts on a CM during rebuild stage. Dust control performance of the rebuilt CM was evaluated using gravimetric sampling and real-time monitoring using multiple personal dust monitors (PDMs). The average results based on about 105 cuts show about 30% and 20% reduction in dust exposure for the miner operator, and the haulage unit operator. Visibility in the mining areas was also significantly improved. The coal company has already decided to implement the system on all their CMs.

Jan 1, 2012

By Sakai Kunito, Nigel B. Sugden

If an Earth Pressure Balance (EPB) TBM is operated effectively tunneling can be carried out producing negligible ground movements. Using the methods given in this paper face losses of the order of 0.5% and less are generally obtainable using an EPB. Operating the EPB TBM in this manner reduces the requirement for the building protection measures used to protect surface structures. This can result in a considerable cost saving for the construction of a tunnel under built up areas. This paper to give guidance on the causes of ground movements around a TBM and the methods of running an EPB TBM that will minimize those movements. Details of how this was successfully done during the construction of the Bangkok Metro are given. The paper is split into three parts as below. 1.Causes of ground movements due to TBM tunneling 2.Methods of controlling those movements using an EPB TBM 3.Case history Bangkok Metro (water main)

Jan 1, 2001

By K. R. Long

Until it was defunded in 1995, the U.S. Bureau of Mines maintained a Cost Estimating System (CES) for prefeasibility-type economic evaluations of mineral deposits and estimating costs at producing and non-producing mines. This system had a significant role in mineral resource assessments to estimate costs of developing and operating known mineral deposits and predicted undiscovered deposits. For legal reasons, the U.S. Geological Survey cannot update and maintain CES. Instead, statistical tools are under development to estimate mining costs from basic properties of mineral deposits such as tonnage, grade, mineralogy, depth, strip ratio, distance from infrastructure, rock strength, and work index. The first step was to re-estimate ?Taylor?s Rule? which relates operating rate to available ore tonnage. The second step was to estimate statistical models of capital and operating costs for open pit porphyry copper mines with flotation concentrators. For a sample of 27 proposed porphyry copper projects, capital costs can be estimated from three variables: mineral processing rate, strip ratio, and distance from nearest railroad before mine construction began. Of all the variables tested, operating costs were found to be significantly correlated only with strip ratio.

Jan 1, 2011

By G. Kordosky, A. Feather, M. Virnig, P. Crane

Introduction The post World War II nuclear arms race and interest in nuclear power generation resulted in a significant demand for uranium in the 1950’s. This led to the development of uranium solvent extraction technology and uranium became the first metal to be recovered on a large scale by solvent extraction (SX) (1), interestingly more than 100 years after the extraction of uranium from an acid nitrate solution using diethyl ether was first reported by Peligot.(2) For a variety of reasons discussed elsewhere (3,4) there have been upheavals in the supply demand picture and price for uranium since the 1960’s. (Figure 1) The low price of uranium from the mid 1980’s until 2004 resulted in a depression of new resource exploration and mine development along with uranium mine / mill closures and decommissioning. People with skills in uranium processing could not find suitable employment and many experienced engineers and metallurgists took jobs elsewhere or retired. This resulted in a significant brain drain with respect to uranium processing technology.

Jan 1, 2007

By Michelle L. Ramos

INTRODUCTION The Ballard Siphon was built in 1935 and consists of two 91.4 cm (36-inch) diameter wood-stave pipes. These pipes convey sewer flowsunder the Ship Canal from northwest Seattle and the Ballard Regulator to the North Interceptor and the West Point Treatment plant. The siphons were constructed using open cut construction, are now buried in the sediment on the floor of the canal, and have severely limited ability to be dewatered, inspected, or cleaned. The existing siphon is 394.7 m (1,295 feet) long from the junction structure on the southern shoreline to the regulator structure on the north. The location of the Ballard Siphon is shown in Figures 1 and 2. The area around the Ballard Siphon has become more populated and developed since the siphon was built. Additionally, the Ballard Siphon is subject to high flows that are greater than the capacity of the pipelines. This has resulted in predict-able Combined Sewer Overflows(CSO?s) whenever a large, long duration rain event occurs. The aged pipeline of uncertain viability has become a critical piece of the sewer system infrastructure. Because of the age of the pipelines and their importance, a sonar survey was completed on both barrels of the Ballard Siphon in December of 2005. The sonar images appeared to indicate crown intrusion of the wood-stave pipes, which suggested that the pipes were in imminent danger of collapse. As a result, an Emergency Declaration was executed by King County and the Ballard Siphon Replacement Project was initiated. Staheli Trenchless Consultants was retained to perform feasibility studies to deter-mine the viability of using microtunneling and horizontal directional drilling (HDD) for replacement of the Ballard Siphon. For the evaluations, the preferred site location was determined to be at or near the site of the existing Ballard Siphon in order to minimize the ancillary sewer work which would accompany moving to a new site. To meet flow demands and hydraulic requirements, the replacement siphon needed a mini-mum equivalent diameter of 182.9 cm (72 inches). This could be accomplished with a number of different siphon diameters and configurations including, but not limited to, three 106.7 cm (42-inch) diameter pipelines, two 152.4 cm (60-inch) pipelines, or two 91.44 cm (36-inch) pipelines and a 121.9 cm (48-inch) pipeline.