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By J. A. Marts, R. C. Gilmore, Jr. Bogin, S. A. Saki, J. F. Brune, J. W. Grubb

The main purpose of mine ventilation design is to provide sufficient quantity and quality of air to the workers and to dilute methane and other contaminants. It is generally perceived that the additional air along the longwall face will improve methane dilution on the face and in the tailgate. However, computational fluid dynamics (CFD) modeling efforts at the Colorado School of Mines (CSM) under a National Institute for Occupational Safety and Health (NIOSH) funded research project have found that higher flow velocities along the longwall face are shown to increase the pressure differential between the gob and longwall face allowing more methane to be entrained in the active face and tailgate (TG) area, thereby negating the dilution effect. The increased face ventilation causes an increase in pressure along the headgate side which allows more oxygen to ingress into the gob area, thereby increasing the amount of oxygen available to form explosive methane air mixtures in the gob and to support spontaneous combustion of the coal. Two things are happening at two different locations along the longwall face, increase in pressure as the flow enters the active area cause an increase in pressure which forces more oxygen into the gob, but as the flow is developed along the longwall face, the higher velocities causes greater pressure drop allowing methane to enter the active face further downstream from the headgate (HG) towards the tailgate side. In this paper, a parametric study will be presented to discuss the effect that varying the face air quantity has on methane concentrations in the tailgate and formation of explosive gas zones (EGZs) in the gob. Counter to conventional wisdom, it appears that increased longwall face air quantities may increase the explosion hazard as they result in increased EGZ volumes in the gob, along with increased methane quantities in the tailgate return.

Jan 1, 2015

By A. Delgado-Jimenez, O. J. Restrepo-Baena

"Sustainability is a concept that involves several aspects like social, economy and environment. For the concept of sustainable mining is an action that involves to do more in social acceptance, distribution of welfare, an stable economy and the environment. Understanding the mining operation that creates a new state that is necessary to prepare and conduct searching for the best results in sustainability. The results of sustainable mining could be directed to achieve less net environmental impacts, more economic development through the transformation of the natural capital and the social acceptance as the result of good negotiations with the communities in the surroundings. To do that, one option is to create frameworks to planning the regional mining developments that allows the stakeholders to decide and develop different strategies to addresses the sustainability in a mining environment. Strategies based on the understanding of the whole mining business, the environment in which it operates, composed for a natural resources network that interact in a complex system, social system composed by different groups, communities and economies that could be subject to big transformations. INTRODUCTION The mining sector represents one of the most significant challenges for sustainable development, given the magnitude of the impacts that it has on the social and natural systems and the potential to generate welfare associated with its exploitation. Some of the main challenges of mining to contribute to sustainable development are the reduction of the environmental and social impacts of mining operations, to advance in the distribution of benefits in the community and the investing in the development of substitutes. To help in this direction it is necessary to integrate more actors than those who have been considered traditionally in the development of the mining business, and consider the natural environment, communities and the several dynamics generated by the interaction of these elements that contribute to sustainable development. This work provides some useful considerations for sustainable development of mining operations in regions with mining vocation, through the convergence of objectives and the finding of synergies between the existing conditions in the territories and the potential conditions that are generated by new mining developments. A DEFINITION FOR SUSTAINABILITY IN MINING The definition of the Brundtland Commission in 1987 established that sustainable development is ""the ability to ensure that it meets the needs of the present generations without compromising the ability of future generations to meet their own needs""."

Jan 1, 2015

By H. Nagel, A. P. Schissler

"INTRODUCTION Anadarko Petroleum Corporation (Anadarko) owns a region of oil shale properties acquired as part of the original Union Pacific Land Grant of the 1800’s in southern Wyoming. These holdings are greater than 2 million acres and occur throughout the Green River Formation in southwestern Wyoming. One area was prioritized for further study by Anadarko Minerals called the Kinney Rim in the Washakie Basin due to low environmental impact, proximity to infrastructure, high ore grade, and low mining cost. The area is a low-overburden oil shale resource that can be developed using surface mining methods, processed by conventional surface retort, and placing waste rock disposal back on the bench for a design capacity of 10,000 barrel per day. The purpose of this papers is to present the results of environmental analysis and timeline, the mine plan, and surface ore processing to offer a resource-ready energy source.Mineral rights for the Kinney Rim area are owned by Anadarko, the United States Bureau of Land Management (BLM) and the State of Wyoming. Mining of oil shale for this study targets Anadarko owned mineral resources.GEOLOGY The study area is in the Green River and Washakie Basins in western Wyoming. The host geologic member is the Laney Oil Shale. Figure 2 illustrates the Laney shale within a generalized lithologic cross section.OIL SHALE LEASING The proposed Kinney Rim oil shale mining project is located in an area dominated by a 1 square mile checkerboard landownership pattern with every other section typically administered by Bureau of Land Management (BLM) and Anadarko. Surface and mineral ownership is a split estate legal entity. The immediate Kinney Rim mine plan is all on Anadarko sections."

Jan 1, 2015

By T. Matthews

"One critical factor in mineral reserves reporting, as well as mine-reserves reconciliation, is the impact of dilution and ore loss in the mining process. This paper considers the implications of various factors on dilution and ore loss projections, by orebody geometry, mining method, and ore control systems. We also discuss the implications of resource estimation methodology on dilution for reserves. DEFINITIONS Dilution and ore loss are two critical concepts in mining, which address the differences encountered between modeled grade and tonnages and actual production. Dilution is volume increase and grade reduction caused by multiple factors, chief among which are mining of waste material along ore-waste contacts, mining of internal waste as ore, and misclassification of waste as ore in the ore control process. There are also geostatistical and estimation aspects sometimes lumped with dilution, including the volume-variance effect (the tendency for small volumes of material to include a less regular distribution of grade than large volumes of material). Ore loss is described as reductions of ore volumes in the mining process due to misclassification of ore as waste in the ore control process, exclusion of discontinuous mineralization from mining due to geometry or continuity, and mechanical losses in the mining process. Dilution is typically taken as an increase in tonnes of ore with no change in contained metal (and thus results in a reduction in grade.) (Some workers consider the likely grade of the diluting material as well, particularly in deposits with gradational grade boundaries—this will be addressed later in the text.) Ore loss is typically taken as a reduction in tonnes at average grade of the ore, and thus results in a reduction in both tonnes and contained metal. If model (resource) tonnes are Mt, grade is Mg, and contained metal is Mm, the resource with ore loss (L) and dilution (D) included (Rt, Rg, and Rm) will be:"

Jan 1, 2015

By K. Lagnese, C. Robinson, S. Muddasani

"Mine water generated from underground coal mining operations contains both dissolved and particulate solids. Dissolved solids primarily consist of sodium, calcium, magnesium, potassium, chlorides and sulfates. When discharged to a receiving stream without treatment, these constituents create a potentially toxic environment for aquatic life. A case study is presented to discuss how a centralized treatment plant treats mine water from six locations to meet discharge limitations for chlorides. Located in West Virginia, this facility consistently achieves less than the NPDES permit limitation of 218 mg/L chlorides in the discharge while generating almost zero liquid waste. The dissolved solids concentration in the influent ranges between 5,000 and 10,000 mg/L, with chloride concentrations of 1,000 to 2,000 mg/L and sulfate concentrations of 2,000 to 6,000 mg/L. The mine water is treated using advanced treatment technology to produce clean water for discharge or reuse. The treatment process comprises aeration, softening, filtration, reverse osmosis, RO reject softening, evaporation, crystallization, final effluent remineralization and sludge dewatering. The solid waste generated in the treatment process is landfilled on site. The leachate is sent back to the facility’s thermal treatment process. Because no liquid waste leaves the property, this plant is termed a “Zero Liquid Waste” (ZLW) facility. INTRODUCTION The advanced water treatment system was built to meet a new 218 mg/l regulatory limitation imposed by the West Virginia Department of Environmental Protection for chlorides discharged to surface waters. The system is designed to treat a maximum flow of approximately 5 million gallons per day (795 m3/h) of mine water. The treatment process utilizes reverse osmosis (RO) membrane treatment to achieve the effluent criteria, and evaporation and crystallization technology to manage the RO concentrate from the water treatment process. As a result, the system produces clean water for discharge while generating zero liquid waste. The effluent water can be used for industrial purposes or discharged to a receiving stream. A Water Impact Index (WIIx) study indicated that discharge of this high-quality water into the Monongahela River watershed may improve the water quality for downstream water users. The waste produced in the treatment process, including softening sludge and salt, is disposed of in an on-site landfill."

Jan 1, 2015

By T. Rockley, B. Goertz, S. McDaniel, J. Brune, K. Bakhsh

"Researchers at the Colorado School of Mines, under a project funded by the Alpha Foundation, have developed a working prototype of a dust sample collection device designed to determine if sufficient rock dust has been placed in an underground coal mine to prevent coal dust explosions. When used in conjunction with the Coal Dust Explosibility Meter (CDEM), the device will provide near instantaneous results on the quality of rock dusting at the tested location. The sampling device applies a pulse of air to a testing surface and collect a representative sample of mine dust based on the pneumatic entrainment process during a mine explosion. The use of both physical testing and Computational Fluid Dynamics (CFD) modeling has allowed researchers to refine the design to accurately represent the particle entrainment and optimize the sample size collected for the CDEM. The prototype is being tested in a mine environment to produce information of usability and practical aspects of mine dust sampling. This paper is a progress report summarizing the current status of sampler development. Researchers expect to have a fully developed and tested sampler in early 2016. INTRODUCTION Underground coal mines face the danger of methane and coal dust explosions. The disaster at Upper Big Branch mine (2010) has demonstrated the impact of a violent coal dust explosion that fatally injured 29 miners. Reviewing the history of coal dust explosions, researchers found that that mine operators and mine inspectors do not have a reliable method of collecting dust samples to assure that sufficient amounts of rock dust have been places to inertize the coal dust. This is a problem specifically with small mines that do not have dedicated safety departments to track rock dusting work. The aim of this research is to develop a simple, portable device that can help underground mine examiners collect dust samples from workings and test the explosibility of the sample near instantaneously. Researchers determined that the current brush-and-pan sampling method required by MSHA (MSHA 2013) is flawed in several aspects: Mechanical sampling with a brush exerts forces on the mine dust that may differ significantly from those present during an explosion. Also, U.S Bureau of Mines research (Nagy 1965) has determined that only the top 0.125 inches of the mine dust layer will be entrained in a coal dust explosion. Using a handheld pan and brush, it is difficult to correctly and repeatably sample a layer of this exact thickness. The following will describe the CFD modeling and experimental testing for the development of the prototype sampler. Researchers analyze the accuracy and performance of the prototype during the course of the build process."

Jan 1, 2015

By Zhan-fang Cao, Hong Zhong, Li-feng Li, Shuai Wang, Xiao-yu Cao, Guang-yi Liu, Ming-ming Wang

"In recent years, there has been an increasing focus on the hydrometallurgical processing of molybdenite, with the decline of high-grade and easily handled molybdenum minerals. In this work, a novel method of selective extraction of complex molybdenite concentrate, which is characteristic of high-grade copper ore, has been investigated. The results show that the leaching of chalcopyrite was impacted by the pH of the electrolyte when sodium carbonate and ammonium acid carbonate were added to the electrolyte. The leaching rate of molybdenum reached 99.39%, but chalcopyrite could not be leached at all in 240 min under experimental conditions of liquid-to-solid ratio 25:1, concentration of NaCl 4 mol/L, pH=9, room temperature and agitation speed 400 rpm. The grade of copper in the residue was 21.84%, with a recovery 99.93%. The technology of selective electro-oxidation leaching of molybdenum and recovery of copper from complex molybdenite concentrate is an effective separation method. The diffusion control model can be used to describe the process of leaching of molybdenum. The activation energy of the leaching reaction was found to be approximately 9.21 kJ/mol.IntroductionWorldwide, high-grade and easy-to-handle ore is decreasing, while the proportion of low-grade and complex ore is increasing. Porphyry copper ores with associated low-grade molybdenite are very abundant in China. Molybdenum is a valuable byproduct in the processing of porphyry copper deposits where the molybdenum is present in low quantities in the form of molybdenite (MoS2). In the conventional processing of copper sulfide ores by froth flotation, most of the molybdenite in the ore floats along with the copper sulfides. In this way, a bulk copper-molybdenum concentrate is produced. Afterward, the bulk coppermolybdenum concentrate is treated by differential flotation to obtain a rougher molybdenum concentrate, which requires several successive cleaning steps to produce a final molybdenum concentrate. However, it is widely recognized that flotation alone cannot economically produce a high-grade molybdenite concentrate, i.e., with less than 0.5% Cu. Therefore, in most cases, it is necessary to use a final step of chemical purification of the molybdenite concentrate to reduce the copper content to a level suitable for the market (Antonijevic and Pacovic,1992; Gerhardt et al., 2001; Fu and Zhong, 2003; Reza et al., 2006, Rafael et al., 2013)."

Jan 1, 2015

By M. Deutsch, E. González, M. Williams

"Uncertainty exists in ultimate-pit limits due to geologic variation and unpredictable economic landscapes. In this work, we show how this uncertainty affects the ultimate pit and how it can be analyzed to improve the mine planning process. A stochastic framework using geostatistical simulation and parametric analysis was used to model the effects of geologic and economic variations on ultimate-pit limits and overall project economics. This analysis was made possible by a new pit optimization implementation that can be automated and set up to calculate ultimate pits for hundreds of different scenarios in a matter of hours. Quantifying ultimate-pit uncertainty early in the mine planning process allows mining engineers to make informed decisions regarding infrastructure placement and to mitigate the possibility of incurring substantial costs to relocate critical mine facilities. Resumen ¦ Existen dudas sobre los límites del pit final por causa de la variación geológica y lo impredecible de los paisajes económicos. En este trabajo les mostramos como esta incertidumbre afecta el pit final y como puede ser analizada para mejorar el proceso de planificación de la mina. Se usó un marco estocástico con simulación geoestadística de análisis paramétrico con el fin de poder hace la modelación de los efectos de las variaciones geológicas y económicas en los límites de pit finales y proyectos económicos en general. Este análisis fue posible gracias a una nueva implementación de optimización de pits que puede ser automatizada y configurada para calcular pits finales para cientos de escenarios diferentes en cuestión de horas. La cuantificación temprana de la falta de certeza en pits finales durante el proceso de planificación permite a los ingenieros de minas tomar decisiones informadas en lo que respecta a la ubicación de infraestructura y en la mitigación de posibles costos sustanciales por causa de la reubicación de instalaciones clave de la mina."

Jan 1, 2015

By William Gleason

"It is safe to say that the coal industry, in the United States and globally, is currently suffering through one of the hardest times in its history.A perfect storm of competition, changing energy demands, politics and public perceptions have collided, inflicting significant harm to the industry that many say is irreparable. This, in turn, has led to massive job losses in some regions, heavy cuts in production and even bankruptcy filings from two of the largest coal mining companies in the nation.Cheaper natural gas can certainly take a large share of the blame for the struggle of coal, as utility companies switch from coal to natural gas for power generation. According to the U.S. Energy Information Administration (EIA) the amount of coal used for electricity generation in the United States is at a 45-year low.November 2015 marked the fifth consecutive month that natural gas was the leading source of power generation in the United States. EIA said generators produced 101,866 thousand megawatt hours (MWh) of electricity with gas in November versus just 87,789 thousand MWh with coal, the lowest monthly level since May 1980 when monthly coal use was 84,884 thousand MWh.According to Robert Godby, director of the Center for Energy Economics and Public Policy at the University of Wyoming, there are a number of factors playing a role in the decline of coal, but he points to the advances in fracking as the tipping point.“A lot of people will argue that regulation is the cause of the decline, and it has played a part, for sure, but the story of the recent past is more about what competition has done to the energy market,” Godby told Mining Engineering. “The war on coal was originally declared by natural gas in 2008.”The price of natural gas, which was at $13 per million Btu before the gas fracking boom, has fallen to as low as $1.68 in December 2015. Likewise, Godby said the cost of wind generation is now lower than the cost of generation from coal and the cost of solar generation is coming down as well. In addition to that, the demand for energy has changed in the past five years with more efficiencies in everything from building design to lightbulbs."

Jan 1, 2015

By K. Hanumantha Rao, A. Javadi Nooshabadi

"Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by chalcopyrite (CuFeS2), which is a copper iron sulfide mineral, during grinding, was investigated. It was observed that chalcopyrite generated H2O2 in pulp liquid during wet grinding and also in the solids when placed in water immediately after dry grinding. The generation of H2O2 during either wet or dry grinding was thought to be due to a reaction between chalcopyrite and water where the mineral surface is catalytically active in producing •OH free radicals by breaking down the water molecule. The effect of pH in the grinding medium or in the water in which solids are added immediately after dry grinding showed that the lower the pH value, the higher the H2O2 generation. When chalcopyrite and pyrite are mixed in different proportions, the formation of H2O2 was seen to increase with increasing pyrite fraction in the mixed composition. The results of H2O2 formation in pulp liquid of chalcopyrite and together with pyrite at different experimental conditions have been explained by Eh-pH diagrams of these minerals. This study highlights the necessity of revisiting the electrochemical and/or galvanic interaction mechanisms between the chalcopyrite and pyrite in terms of their flotation behavior.IntroductionHydrogen peroxide causes nonselective oxidation of sulfide minerals if present in pulp liquid. Oxidation of sulfide minerals takes place when they are exposed to atmosphere and during the grinding process when the particle size is reduced for flotation. The oxidation of chalcopyrite leads to iron dissolution from the particle surface and from a metal-deficient, sulfur-rich surface (Fairthorne et al., 1997). Other studies report the formation of iron hydroxide on the surface with the metal ion dislodged on the surface (Buckley and Woods, 1984; Chander, 1991). The chemical reactions for the oxidation of chalcopyrite in alkaline and in acidic solutions with the addition of oxidants, such as oxygen or hydrogen peroxide, are shown in Eqs. (1) and (2), respectively (Shi and Fornasiero, 2010)."

Jan 1, 2015

By M. L. Free, N. Dhawan, M. S. Safarzadeh, J. D. Miller, M. S. Moats

"The drum agglomeration of nickel laterite and gold ores has been optimized through the design of experiments (DOE) using a Taguchi L16 (45) orthogonal array to determine the optimum conditions for maximizing average agglomerate size and minimizing the amount of fines. The effects of controllable operating factors including moisture content (nickel laterite ore: 34-37%; gold ore: 7-10%), retention time (2-3.5 min), drum speed (15-45% critical speed), drum load (nickel laterite ore: 8-32 %; gold ore: 6-22%) and acid concentration (150-600 g/L) on the performance of the agglomeration process were studied.For nickel laterite ore, maximum average agglomerate size and minimum percent fines (-1 mm) occurred under the following conditions: drum load (23.7%), moisture (36.5%), time (3 min), drum speed (30% critical speed) and acid concentration (150 g/L). Under the studied nickel laterite ore conditions, the most effective parameters for maximizing average agglomerate size and minimizing the amount of fines were found to be drum load and acid concentration, respectively. Drum speed had a statistically significant effect on minimizing the amount of fines. Maximum average agglomerate size and minimum percent fines (-1 mm) for gold ore occurred under the following conditions: drum load (19.3%), moisture (8.5%), time (2 min 15 s) and drum speed (40% critical). The most significant factors for maximizing average agglomerate size and minimizing the amount of fines for gold ore were found to be drum load, time and moisture.IntroductionSeveral heap leach operations have experienced problems in terms of poor recovery due to percolation issues caused by low-grade complex ores, tailings and clayey deposits. Poor percolation can lead to low metal extraction due to solution channeling or the development of impermeable/dead zones within the heap (Dhawan et al., 2012; Kappes, 2005). Improper heap building practices such as the use of loose agglomerates and inadequate attention toward the presence of clay minerals have been among the main reasons for percolation issues. During the transport of ore material, severe segregation of the material can occur. To overcome percolation problems, a major improvement was made through the introduction of agglomeration prior to ore placement. If the ore particles and agglomerates are of similar size, segregation can be avoided to a great extent (Dhawan et al., 2013; Kinard and Schweizer, 1987; McClelland and van Zyl, 1988). Agglomeration improves the uniform percolation of solution through ore heaps and is applicable to many ores, wastes and milled tailings (Bouffard, 2005; Dhawan et al., 2013). Manning and Kappes (2005) reported agglomeration/ stacking accounts for ~14% of the total heap leaching operating cost. The cost of binder is a primary contributor toward the total cost. Also, the lack of consistent quality control tests for agglomerate often leads to operational problems. The details are mentioned elsewhere (Dhawan et al., 2013). Considering the significant contribution of the agglomeration/stacking step toward the total heap leach operating cost, it is appropriate to study the agglomeration behavior and its optimization for two different ores. The reason to study the same parameters for different ores is the fact that significant variability in the agglomeration behavior of different ores has been reported to exist in industrial operations."

Jan 1, 2015

By L. Giraldo, S. Cotten

"INTRODUCTION The vision for a “Mine Escape Vehicle” (MEV) is to provide capabilities to better enable underground travel to allow miners to escape to safety in the event of a major mine emergency such as a fire or explosion. Enhanced transportation and survival capability such as a MEV may have minimized or eliminated the loss of life associated with several of the mine disasters that occurred in 2006, which together resulted in the deaths of 19 miners. Although the most desirable response in an emergency is for the miners exit the mine, if escape proved to be impossible, the miners could use the vehicle to proceed to an underground point of refuge to await rescue. Under a NIOSH contract issued in April 2008, Raytheon developed a conceptual vehicle design including (1) the capacity to provide life support functions for a number of miners, (2) operate in a reduced oxygen atmosphere with little or no visibility, and (3) travel at speeds faster than miners could walk out of the mine. The proposed MEV concept envisioned modification of a personnel carrier (a.k.a. a “mantrip”) that would be used every day during normal mine operations, not a specialized vehicle reserved only for or dedicated solely to escape. The MEV concept also included escape aids that are detachable from the vehicle in the event that the vehicle must be abandoned on the way out of the mine and the dismounted miners had to proceed on foot. To advance MEV development, NIOSH awarded a second contract in 2010 to Raytheon to build, test, and evaluate the performance of a proof-of-concept MEV that culminated in a 7-month performance evaluation of vehicle in an Eastern Kentucky coal mine (Figure 1). User feedback and lessons learned from the proof-of-concept vehicle evaluation provided support for the general MEV approach, guidance related to the practicality of MEV capabilities, and indication of specific technology shortcomings for this application."

Jan 1, 2015

By M. Zaske, E. R. Mortensen

"As mine operators continuously look for ways to improve functionality and increase the capacity of their facilities, existing process equipment is often pushed to and past its designed limitations. As a result, equipment manufacturers are asked for economical solutions to meet redefined processing needs. Newmont Mining in Nevada was faced with such a need in its carbon-in-leach (CIL) train. In a collaborative effort between Newmont and Derrick Corp., tests were conceived to assess the viability of drop-in replacement screens using a three-dimensional screen panel. The synthetic pyramid screen panel was designed to increase flux while maintaining the existing equipment footprint. The test is the end result of an ongoing research and development project that Derrick undertook in an effort to refine and upgrade its urethane screening media. The application of this new screening media is presented for this CIL application along with future plans to implement the technology in trash, safety, carbon dewatering, carbon sizing and other scalping applications.In spring 2014, Process Machinery Associates (PMA), Derrick’s local area representative, met with Newmont to discuss the operation of the in tank screens in the CIL train at Twin Creeks, near Golconda, NV. The process was originally designed for about 12+ kt/d (13,000+ stpd) but the throughput has been steadily rising since the plant was commissioned. Newmont now requires 14.5 kt/d (16,000 stpd) to be processed. The original screens were running at their maximum capacity, and Newmont required additional screening capacity to continue the upward trend in production.Physical separation by use of screening is one of mineral processing’s oldest methods. Screening media can be prone to blinding and pegging, and the necessary support structures required for finer apertures result in smaller open areas. Therefore, more screening area generally requires larger panels and, thus, larger machines, screening media that does not blind or a combination of both.In an effort to avoid larger machines, Derrick had been working on a revolutionary threedimensional panel that would add nonblinding screen area in the same footprint of an existing machine. Alpha testing of the new screen panels began at Newmont Mill 2/5 in July of 2013. It was still occurring when PMA approached Twin Creeks about beta testing a refined machine and pyramid panel. After presenting the initial results, Twin Creeks determined that an upgraded machine and pyramid panels would be a possible solution to meet production needs. The testing of a machine eliminated the risk of capital costs to Newmont and provided Derrick the opportunity to obtain additional data."

Jan 1, 2015

"Above ground, elevated belt conveyors with span >200m between support points can cross valleys, rivers, forests and buildings with minimal footprint. The Cable Bridge Conveyor is a new type of suspension bridge based belt conveyor system that has long span between support points. It combines typical belt conveyor and the improved simple suspension bridge to form a conveyor bridge. The improved simple suspension bridge places a widened bridge deck directly on suspended cables, without the need for high truss towers. The benefits include reduced cost and increase aerodynamic stability compared to conventional suspension bridges. The Cable Bridge Conveyor’s compatibility with current belt conveyors allows the versatility in application. INTRODUCTION Belt conveyor is a ubiquitous material handling system. Conveyors can be placed directly on ground or elevated above ground by support structures. Considerations of cost, design, environment, civil work, safety and process requirements are all factors in choosing the ground based or elevated conveyors (1). The ground based conveyor has the lowest structural cost, where the conveyor is laying on an existing road. Maintenance access is easy and straight forward. Compared to the ground based conveyor, the elevated conveyor can cross existing buildings and structures, give passage to traffic beneath the conveyor, has smaller footprint, and requires less civil work to pass mountainous areas. The drawback of the elevated conveyor is higher structural cost, which can be compared with savings in civil work and other benefits. Conventional elevated conveyors are supported by rigid truss type structures, with typical span of 15m to 25m between concrete support bents or steel trestles. Triangular truss, using steel pipes instead of hot rolled steel sections, is more efficient. The typical span of triangular truss supported conveyor is around 30m to 40m, while the weight is about 10% less than conventional box truss with shorter span (2). Long span elevated structure based on suspension bridge is another step forward in increasing the distance between support bents, reaching 100m~400m and beyond. The very wide distance between support points of an elevated conveyor offers unique benefits. The conveyor can cross rivers, valleys, highways, forests and other obstacles with minimal footprint and reduced civil work. The versatility from long span conveyors makes bulk material transportation over previously impossible terrains now a feasible option."

Jan 1, 2015

By R. A. Thomas, C. D. Litton

"The objective of this study was to develop a laboratory-scale method to rank the ignition and fire hazards of commonly used underground mine materials and to eliminate the need for the expensive large-scale tests that are currently being used. A radiant-panel apparatus was used to determine the materials’ relevant thermal characteristics: time to ignition, critical heat flux for ignition, heat of gasification, and mass-loss rate. Three thermal parameters, TRP, TP1 and TP4, were derived from the data, then developed and subsequently used to rank the combined ignition and fire hazards of the combustible materials from low hazard to high hazard. The results compared favorably with the thermal and ignition hazards of similar materials reported in the literature and support this approach as a simpler one for quantifying these combustible hazards. Resumen ¦ El objetivo de este estudio fue el desarrollo de un método a escala de laboratorio con el fin de clasificar los peligros de ignición y fuego de materiales comúnmente usados en minería subterránea y la eliminación las pruebas costosas y a gran escala que se realizan actualmente. Se usó un aparato de panel radiante con el fin de determinar las características térmicas relevantes de los materiales: tiempo de ignición, flujo de calor crítico para ignición, calor de gasificación y tasa de pérdida de masa. Se derivaron tres parámetros térmicos de los datos: TRP, TP1 y TP4, luego fueron desarrollados y subsecuentemente usados para clasificar los peligros combinados de ignición y fuego de materiales combustibles, de bajo riesgo a alto riesgo. Los resultados tuvieron una comparación favorable con los riesgo/peligro térmicos y de ignición de materiales similares reportados en la literatura y apoyan este enfoque como uno más simple para la cuantificación de estos riesgos combustibles"

Jan 1, 2015

By C. Dirscherl

"Declining ore grade, leading to increased material transport and plant throughput requirements, is one of the major challenges in the current mining environment. This is accompanied by rising energy and labor costs, decreasing plant productivity and the requirement for highest plant availability. Siemens’ answer is the latest gearless drive technology for grinding and conveying. Gearless drive solutions are surpassing the mechanical limits of conventional drive systems serving to utilize the principle of economies of scale. By elimination of various components of the drive train, the maintenance activities and spare parts inventories are reduced. The advantages are culminating in the incomparable availability of the Siemens’ gearless drive technology of 99.46 to 99.71 percent — as confirmed by real operational data from the Antapaccay Mine in Peru. It was the first plant worldwide to use gearless drives for both applications.The gearless principleThe principal idea behind a gearless drive system is to reduce the number of components and diversity of parts, especially those components that are at the edge of current manufacturing or power limits are eliminated. The used equipment provides the possibility to scale up and increase the power transmitted to the driven machine. This finally results in a possible increase of throughput respectively transported or processed material.In order to apply this principle of simplification, the components of a conventional drive system are examined. They consist of:• Driven equipment (mill, conveyor, slurry pump).• Components altering the rated speed (gearboxes, girth gear/pinion).• The electric motor.• Starting devices for the motor (VSD, CCV, LRS, clutch), which are necessary because a direct across-the-line start of high power motors with load would result in a significant drop in the line supply voltage.• Alternatively a variable speed drive (VSD) or cycloconverter (CCV) to utilize the possibility of speed variation during operation.• The electrical supply system (cables, switchgear, transformers, protection equipment)."

Jan 1, 2015

By S. Pastor, S. Hahn

"As the concept of autonomous haulage systems (AHS) moves from prototypes to production-ready applications, the operating performance of the haul road will become critical to the overall success of autonomy in mining. Deterioration in road performance will require remediation, human intervention and significant, albeit temporary, changes to operating procedures, to accommodate road maintenance. With autonomous trucks comes the opportunity to reduce road formation and operating width. This will, therefore, generate potentially significant reductions in stripping ratios and improvements to mine economics, but only if the design of the road, and its associated deterioration rate, is predictable and manageable, based on the materials used to construct the road and the maintenance interventions to be carried out on the road.This article presents the assessment of a digital photogrammetric approach to the characterization of a mine haul road surfacing, or wearing course material. Design and maintenance parameters for these roads are currently determined partly from physical particle size distribution (PSD) assessments of ‘in-road’ material. However, this approach does not accommodate the need, nor leverage the benefits from, AHS. The digital imaging approach is assessed in this article, using two standard digital analysis software tools, together with in-service mine haul road wearing course samples and images that are analyzed to determine the applicability and reliability of the approach to deliver a real-time assessment of haul road surfacing conditions and thus to identify the road maintenance interventions required to address deterioration.Enhancements in mine road design, construction and management for the past 20 or more years has largely been focused on short-term tactical needs. These improvements have not been matched by advances in the fundamental technologies that underpin mine road design and management and, although the quality of mine roads have probably"

Jan 1, 2015

By G. Plenge, M. Brittan

"Plant design parameters for gold extraction, leach residence time and cyanide consumption are generally determined from standard bench-scale bottle roll or agitation leach tests. The application of laboratory data to process design has essentially evolved from the testing of oxide or low cyanide-consuming ores. Such scale-up factors may not be appropriate for ores that deplete cyanide and oxygen since this interferes with gold extraction kinetics. High cyanide-consuming ores also exacerbate the differences between the conditions applicable to small batch tests and large continuous operations. Testing a reactive high-sulfide South American ore deposit necessitated the re-evaluation of conventional bench-scale cyanidation procedures. Pulp density can be a major factor affecting the results for plant design purposes and should be treated as a test work variable. As with high clay-bearing ore, where pulp viscosity can hinder mass transfer, minimizing the influence of test equipment on gold extraction kinetics may require low pulp densities. This introduces interpretative complexity since cyanide consumption changes as a function of pulp density. Activated carbon also consumes cyanide, so its effect has to be accounted for in the case of carbon-in-leach (CIL) or carbon-in-pulp (CIP) plant flowsheets. Since cyanide is an expensive reagent, often a major component of plant operating costs, it is important that the scale-up of cyanide consumption from laboratory to plant be representative. The traditional method of reporting cyanide consumption in bench-scale testing is questioned since it takes credit for residual free cyanide in leach tails. The projected plant consumption may be better estimated by using the total cyanide addition in a laboratory test. We propose a test work protocol that categorizes the cyanide consumption into its components but requires more extensive testing to generate design data. The ultimate validation of scale-up factors used by project engineers requires rigorous bench-scale versus commercial plant comparative data. Such data are not readily available.IntroductionMajor factors in the design of a gold processing plant are gold recovery, leach residence time and cyanide consumption. These parameters are important determinants of project revenue and costs and, thereby, potentially of project financial viability. Estimates of the associated design criteria for a commercial plant are normally generated from laboratory cyanidation, carbon-in-leach (CIL) or carbon-in-pulp (CIP) tests using standard procedures (Angove, 2005). The key question is whether the standard laboratory test work procedures developed historically for low reagent consuming ores are appropriate for ores that are highcyanide and oxygen consumers.The present paper examines test procedural factors that can influence the test work results and, therefore, their interpretation for design scale-up purposes. The study refers in particular to a high cyanide-consuming sulfide ore. The ore was being tested to develop scoping level engineering design criteria for a proposed project in South America."

Jan 1, 2015

By P. Schumacher, M. Fein

"INTRODUCTION Newmont’s Twin Creeks operation, part of its Nevada operations, produces gold from two large open pit mines. Mined ore is processed at one of three on-site facilities: a refractory mill with two autoclaves, an oxide mill and heap leaching facilities. Twin Creeks also processes various off-site sourced concentrates and ore including high grade underground ore from the Turquoise Ridge joint venture with Barrick Gold Corporation. This paper discusses Twin Creeks’ strategic mine planning process which is part of an in-depth continuous improvement program called “Full Potential.” The process involves cost analysis, phase design validations, life-of-mine production scheduling optimization, mill blending strategy and stockpile management. Through this Full Potential strategic planning, Twin Creeks is creating significant value by increasing near term free cash flow and life of mine net present value (“NPV”) and achieving cost reductions. TWIN CREEKS OPERATION Twin Creeks is one of the major gold producers in Newmont’s Nevada operations (see Figure 1, 2 and 3). Located approximately 15 miles north of Golconda, Twin Creeks produced approximately 28 % of the Nevada operation’s 2013 gold production of 1.8 million ounces (Newmont 2014). The production included Turquoise Ridge sourced gold ore from the joint venture (75% Barrick Gold Corporation, 25% Newmont).At Twin Creeks, there are two large open pits, named Mega Pit and Vista Pit (see Figure 4, 5 and 6). There are 3 to 5 phases in Mega Pit and 2 to 5 phases in Vista Pit depending on the gold price assumption. There is also an underground exploration project at the bottom of the Vista Pit. Twin Creeks produces gold from a variety of ore types requiring different processing techniques depending on economic and metallurgical characteristics. To ensure the best use of processing assets, engineering and processing groups at Twin Creeks utilize a Mixed Integer Linear Programming (“MILP”) model to guide the flow of mine sequences and material routing to various process facilities (Kawahata, Schumacher and Hufford 2013)."

Jan 1, 2015

By J. L. Ellenberger, G. S. Esterhuizen, T. Miller, M. M. Murphy

A limestone mine in Ohio has had roof instability problems that have led to massive roof falls extending to the surface. This paper describes a collaborative study of these events undertaken by the National Institute for Occupational Safety and Health and the mine owners. The study focuses on the roles that weak, moisture-sensitive floor and weak bands in pillars play in a general instability that led to roof falls at this mine. The NIOSH Stone Mine Pillar Design software (S-Pillar) does not include analysis for weak floor or weak bands and recommends that when such issues arise they should be investigated further using a more advanced analysis. It is important to recognize that without the impact of the floor or weak bands, the calculated inherent stability factor for these pillars is within the recommended design range. Therefore, to further investigate the observed instability occurring on a large scale at the Ohio mine, FLAC3D numerical models were employed to help determine the failure mechanism. Roof-to-floor convergence monitors were also installed in the areas predicted to have further instability. This case study will provide important information to limestone mine operators regarding the impact of weak floor causing the potential for roof collapse, pillar failure, and subsequent subsidence of the ground surface. This recent case demonstrates that weak floor can be a significant factor in pillar stability and should be assessed during the design stage of the mine.

Jan 1, 2015