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By C. St. J. Lowe

Weir Minerals ? Linatex has developed a new approach to the dewatering of fine solids, particularly mine and process plant tailings and wastes. The Weir process, patented in 2009, uses standard process equipment in a standard flow sheet, but applies this approach to materials previously thought to be too fine to be dewatered with this type of equipment. The process can be applied to a wide range of materials, making lower moisture products that can be re-used elsewhere or sent to disposal. It has already been tested successfully on a small scale in Australia on coal fines, zinc mine tailings, red mud, iron ore tails, and copper/gold tailings. The principal factors affecting the solids recovery have been established as: 1. The % solids in the slurry feed to the process ? a higher % solids provides a higher recovery. 2. The particle size of the feed solids ? coarser feeds provide higher recoveries. The key factors affecting the % solids in the dewatered product have been determined to be: 1. The particle size of the dewatered product - coarser materials produce a higher % solids in the dewatered product. 2. The height of the discharge weir on the dewatering screen - a higher discharge weir on the dewatering screen also yields a higher % solids. 3. The angle of the dewatering screen. Setting the screen horizontal increases the travel rate, reduces the bed depth and reduces the % solids in the dewatered product. Back-sloping the screen up to 5° reduces the travel rate and increases the % solids in product. Eight plants ranging in size from 3 t/h to 40 t/h have been sold to a range of customers in China, for coal and iron ore tailings treatment. Australia has one plant, a small unit rated at 30 t/h of sand fines, and has recently completed test work on iron ore tailings for a dewatering plant rated at 135 t/h. Additional tests are under way with clients around Australia, to demonstrate the capabilities of this process on specific applications and to provide confirmation that materials can be processed successfully, before moving on to larger scale tests or full size plants. The advantages of the Weir process include - lower moisture in products, uses proven equipment, lower capital and operating costs, lower maintenance costs, reduced power consumption, compact footprint, can be used in parallel or in series with existing equipment, modular design to facilitate redundancy or expansion, efficient water recovery and simpler water management, reduced cement in backfill mixes, smaller tailings disposal footprint, reduced earthworks requirements, much lower seepage losses, more stable tailings mass, progressive site rehabilitation, reduced impact on local wildlife, reduced risk of dam wall failure, and reduced environmental monitoring and compliance costs.

Feb 23, 2014

By K. V. Raj, T. Bhowmick

"Control of fugitive dust in high latitude open-pit mines is challenging due to unique atmospheric phenomenon resulting in complicated flow regimes, and atmospheric inversion due to lack of adequate solar insolation during prolonged winter seasons. In this study, two idealized (One conical and one trapezoidal) three-dimensional open-pit mine geometries are simulated for differing seasonal conditions using the CRADLE-CFD package. The airflow is modeled using Standard Kappa-Epsilon Turbulence Model solved by RANS equations. The concept of particle tracking is used to predict the flow pattern of dust particles. For various climatic conditions, and two different pit geometries, fugitive dust particles varying in size (PM0.1 to PM10) and concentration are generated at various locations of a mine and dispersed by the airflow pattern inside the open-pit mine. The amount and location of dust particles inside the pit is reported at various time intervals. The development of atmospheric inversion during the winter season significantly affected the amount of dust retention inside the open-pit domain. While during the summer, airborne dust particles were transported outside the open-pit domain in a short time. INTRODUCTIONComputational fluid dynamics (CFD) modeling of open-pit mines is challenging. In a recent paper, Bhowmick et. al. [1] presented different constraints and challenges in developing a 3-Dimensional CFD model. In order to obtain realistic results from a simulation, however, appropriate values of various inputs are required. Simulation results of a 3D non-hydrostatic model by Shi et al. [2] showed that both the mechanical and thermal forcing are important mechanisms, controlling the development of the atmosphere inside an open-pit. Alvarez et al [3] used a CFD model for predicting dust dispersions in a medium sized limestone quarry, and presented several iso-surfaces of varying dust concentrations using ANSYS CFX. A similar study by Torno et al [4] introduced a more sophisticated parameterization of the dust sources. Flores et al. [5] used an OpenFOAM CFD simulation package to simulate and predict pollutant dispersion in an idealized and an actual open-pit mine in Chile with an intense insolation condition. The model was simulated for three different conditions: (i) air advection (10 m/s) driven mechanical turbulence; (ii) surface heat flux (240 W/m2) driven thermal buoyancy; and (iii) both air advection and surface heat flux driven conditions. The results of the simulations showed that the buoyant currents contribute to the removal of a large percentage of the dust particles. The buoyant currents modify the purely mechanical-turbulence induced flow pattern and reduce the particle residence time, commonly observed in a purely mechanical turbulent flow (non-buoyant case)"

Jan 1, 2015

By Felicia F. Peng

One of the serious environmental problems related with the extraction of coal for carbon and energy resources is that upon preparation of coal a large quantity of fine coal has been discarded and/or lost to tailing impoundments. These problems present hazardous conditions to the local community. Also, the precious heating value of energy resources has been wasted and Left unrecovered in the tailing impoundments. Both RCR and SMRC ACTs recognize and mandate an investigation into slurry impoundment problems, including consideration of potential use of those coal fines in the tailing ponds. Bituminous coal producing district 46 encompasses the northern panhandle of West Virginia. The major producing seam in this district is the Pittsburgh coal, 1-7 meters thick. 75 percent of the coal producted in this district is deep mine. This district produces approximately 3 million tons of coal waste materials annually. One of the coal preparation plants in the district which we have visited has been processing Pittsburgh too. 8 seam coal for the past ten years. The characteristic data of the Pittsburgh raw coal is shown in Table 1. The raw coal produced in this district is characterized by its high heating value, high ash and high sulfur contents. The coal preparation plant flowsheet designed and used to process this Pittsburgh raw coal prior to 1979 is shown in Figure 1. It was a partial coal cleaning plant' with 1 092 mt/hr designed capacity, which consisted mainly of a coarse coal cleaning circuit. The rejected fine particles in the size range of -600µm which consisted of vorsiv underflow and cyclone underflow from clean coal streams and spiral classifier overflow from refuse stream were settled down in a thickener with the aid of anionic flocculant prior to being pumped to the tailing ponds for disposal. Figure 2 shows that the same coal preparation plant in Figure 1 was modified to a semi-complete coal cleaning plant and expanded its designed capacity to 1 820 mt/hr, when market of coal fines becomes feasible in 1979. Tabling in the clean coal circuit and thermal dryer in the dewatering circuit were added to the flow sheet. The purpose is to recover more coal fines, reduce more pyritic sulfur and ash contents in coal and also to reduce solid waste material to be stored into the slurry impoundment, giving them a Longer Life. The rejected fine particles fed to the thickener are clean coal fines from two cyclone overflows and the fine refuse from the rufuse cyclone overflow. The clean coal fines presently lost to the tailing impoundments are in the size range of -150 µm and constitute about 5 percent of the overall plant feed. Total amount of tailing waste material accumulated in a tailing impoundment is estimated at 1.3 million metricton for the past ten years.

Jan 1, 1982

By D. Kruse, M. Sanaei

"A three dimensional nonlinear finite element model was constructed to simulate and analyze the behavior of large diameter segmentally lined shafts when subjected to seismic waves. To reduce computation time, a less complex shell model was also constructed. Transient excitation was applied to the supported nodes of both models, at the base of the shaft, with a time-function representing the 1994 Northridge earthquake. The models are developed at the end of an original phased construction-stage analysis of the shaft. Displacements, forces and moments are monitored and compared during the analyses steps. During the initial 15 seconds of the earthquake, the top and bottom of the shaft show similar displacements. The maximum horizontal displacement within the shaft is approximately 3.5 inches in the solid model and the shell model results conform to this by a factor of 50%. Keywords: Segmentally Lined Shaft, Soil Structure Interaction, Seismic Analysis, Finite Element Modeling, Ground Modification INTRODUCTION TO SEGMENTALLY LINED SHAFTS Large diameter segmentally lined shafts (SLS) are underground systems, constructed sequentially using precast concrete segments to form a ring which provides both temporary earth support during excavation and permanent structural support. The segments are connected via alignment pins, bolts, packer material and other structural components that ensure the continuity of the SLS. In this analysis, the construction process uses a top-down sequence wherein each soil layer is excavated, damp-proofing is installed and concrete segments are assembled to form each ring. This process is repeated until design depths are achieved completing the shaft. It is known that cylindrical structures sustain lateral (earth) pressure mainly through the mechanism of hoop-stress. As Timoshenko’s ring theory goes, a cylinder shell structure is prone to buckling under external pressure [1]. In typical designs of small diameter shafts and tunnel linings with small diameter to wall thickness ratios, nonlinear buckling analysis (i.e., nonlinear geometric analysis) is not considered, since the buckling capacities for those structures are much higher than typical pressures encountered at their design depths. However, in shafts with large diameter to wall thickness ratios, buckling could become a critical issue. In any given mode, as buckling occurs in a ring, some sections of the structure move inward while alternating sections protrude. Although the presence of the surrounding soil could mitigate the buckling by resisting the movement (Soil structure interaction) seismic loads can cause ground movements and lateral pressures, exceeding the capacity of the structure. It is therefore essential to design the load bearing capacities of the shaft to minimize the possibility of any catastrophic failure due to earthquakes. This study explores the structural response of the SLS when subjected to seismic loads."

Jan 1, 2016

By W. L. Wilson

This article could be entitled "The Rediscovery of the Windfall Mine," since it was originally discovered early in this century. The "Rediscovery" was made by personnel of Idaho Mining Corp. during the course of an investigation of the Windfall mine beginning in 1967. There had been historic production from the Windfall mine sporadically from 1908 until 1951 and there was present during much of this time a mill that vat leached crushed ore. Mining was underground to a depth of 91 m (300 ft). Prior to Idaho's activities, the mine had been inactive since the early 1950s. Idaho Mining Corp. was an "exploration company," privately owned and based in Reno, NV. The writer was at that time manager of exploration for Idaho Mining. The company held interests in the Cortez Gold mine, then being readied for production, and had conducted exploration for various mineral commodities in Nevada for several years. Idaho's attention was first directed to the Windfall mine in the Eureka District, Eureka County, Nevada (see Fig. 1, Nevada map) by a search of published geologic literature of the state of Nevada. An attempt was made to acquire a library consisting of all of the published geo-data for the state of Nevada as well as to obtain as much of the unpublished material as could be acquired. A study of these data, and in particular Tom Nolan's Professional Paper 406 (1962), indicated that the Windfall mine was anomalous for the Eureka District in that it contained deposits of disseminated gold. This source and other sources indicated that the Windfall had produced from underground operations approximately 59 kt (65,000 tons) grading 12.6 g/t (0.368 oz per ton) gold ore that was processed in a vat leaching operation and that the gold was "invisible" to the naked eye and could not be panned. It was ascertained that there were no active operations near the mine and a geologic reconnaissance was made and a few samples were taken to attempt to determine if the mineralization was indeed disseminated. The initial group of samples around the mine area averaged 1.25 ppm [11.5 g/t gold (0.036 oz per ton)] and the mineralization did appear to disseminate into the wallrock. The next step involved researching the ownership of property in the immediate area of the mine, and [ ] a property map was constructed that showed six or seven different ownerships of patented and unpatented claims, as well as some open ground, in the target area. Negotiations were commenced with the claim owners, ultimately resulting in the acquisition of all of the desired property, although some of

Jan 1, 1990

By Michelle Mangione, John Kuyt, Zili Li, Michael Mooney

"Twin bored tunnels are commonly used for a variety of road and tunnel projects throughout the world.A major component of these twin tunnel projects is the construction of connecting cross passages that provide emergency access between tunnels and space for tunnel service equipment. Cross passages can present significant geotechnical risk and technical challenges to a project – support systems must be considered for redistributing loads around the mainline tunnel openings and supporting new loading resulting from excavation activities. Additionally, cross passages pose risks to construction schedule and cost because of their sequencing at the end of construction, so mitigation of schedule overruns can be critical. By observing field instrumentation data records from completed projects, a better understanding of the force development in each temporary structure can be achieved with indication of which structures are the most critical to support ground loads.This article discusses the data obtained during the Brisbane Airport Link construction project, a road tunnel located in Brisbane, Australia. Strain gauge instrumentation data from the precast segmental lining and propped opening steel sets is presented with relation to the construction sequence and local geology. Interpretations and conclusions based on the observed trends are proposed, with several key mechanisms affecting the loading and unloading identified.Brisbane Airport Link project overview. The Brisbane Airport Link project in Brisbane, Australia, connects central Brisbane to the airport located northeast of the downtown with more than 7 km (4 miles) of twin-bore road tunnels, two lanes each, with several caverns, interchanges and access ramps constructed along the alignment to connect to other key routes in the city. The project was executed under a public-private partnership (P3) scheme, funded and operated by the consortium BrisConnections. The main contractors were a joint venture of John Holland and Theiss, while the design was carried out by Arup and Parsons Brinckerhoff.Construction of the tunnels was conducted using a combination of 17 roadheaders and two 12.5-m (41-ft) double-shield Herrenknecht tunnel boring machines (TBM) (Harding & Francis, 2013). The TBMs were used in earth pressure balance- (EPB) type closed mode for the initial portion of their drive, starting at the northeast end of the alignment. Tunneling started in soft ground, weathered rock, and mixed face conditions initially, progressing to competent sedimentary rocks as the drive continued. Support was accomplished using a 12.1-m (39.8-ft) diameter steel fiber reinforced concrete (SFRC) segmental lining installed behind the TBMs in nine segments + one key arrangement. The innovative design of these segments is detailed in the paper by Harding and Francis (Harding & Francis, 2013). Ancillary roads and underground interchange caverns were constructed by sequential excavation techniques using roadheaders, accounting for nearly half of the total underground excavations."

Mar 1, 2017

By Leonard G. Austin

INTRODUCTION Run-of-mine (ROM) coal generally falls into two major groups: that from underground mining (continuous mining machines) and that from strip mining. Continuous miners produce a finer product and strip mining a coarser product that is crushed to produce the desired size. Further size reduction depends on the proposed use or uses for the coal. For example, for pulverized coal power plants as a rough general rule, the coal is ground or pulverized to 80% less than 200 mesh (0.075 mm) before it is blown into the boiler furnace as a coal-air suspension. For solvent-refined coal processes, the desired size is about <30 mesh (< 0.60 mm); in fluidized bed uses, the coal is typically < 8, in. (< 1.6 mm); for Lurgi pressure gasifiers, the desired size is 1 to 2% in. (25 to 63 mm). Coal preparation plants typically work with a coarse size fraction of 6X 1/14 in. (150X32 mm). The past decades has shown a steady increase in the tonnage of coal cleaned, 1 but it seems probable that the next decades will produce much wider applica- tion of deep coal cleaning before use and that the stages of crushing and pulverization from ROM coal to end use will be integrated with coal cleaning processes. Since different cleaning processes work better with different ranges of coal size and the degree of cleaning possible is intimately connected with the size of coal, a systems design engineer will need as much information as possible concerning the size distributions produced by different types of machines and how these can be modified by different ways of operating the machines. In general, the following questions will be of concern: 1.What type of crusher or grinder is best, or most satisfactory, for reducing a given feed size and type of coal to a desired product size? Frequently, the desired product is smaller than some nominal size, but with as few fines as possible, or the fraction in some size range has to be as high as possible. 2. How big does the machine have to be for a given throughput rate? 3. How much electrical energy (or its equivalent) is required per ton of product? 4. How does the size distribution vary with change in throughput rate, and is a control scheme necessary or desirable? 5. Can the size distribution be readily varied if desired? 6. What is the optimum way to operate a machine in a given system? This section will emphasize certain general principles which guide us in answering these questions. PARTICLE SIZING Since particle size and size distribution are the raw information with which we work, it is worthwhile to review the terminology used. In this regard, see

Jan 1, 1991

By M. O. Woodburne

The Mojave Desert province is comprised of a diverse group of mountains that range in elevation from about 2,000&apos; (600 m) to 3,000&apos; (900 m), separated by tracts of alluvial cover. The province is bounded on the southwest by the San Andreas fault, on the north by the Garlock fault, on the east by the Death Valley- Granite Mountains fault. Rocks preserved in the Mojave Desert Province collectively record a rich assemblage of Precambrian crystalline basement&apos; ranging in age from ca. 1.8 to 1.35 Ga, upper Precambrian and Paleozoic Cordilleran miogeoclinal sediments and platform rocks, Mesozoic backarc and interarc shallow marine and continental sequences, locally emplaced Permian- Triassic syenitic plutons, extruded hypabyssal volcanic rocks of the ?Jurassic "Sidewinder Volcanic Series," and regionally intruded late Mesozoic granitoid batholithic rocks (Dokka et al., 1988). The Mojave block apparently was a high-standing element for much of the early Tertiary, with strata of Late Cretaceous or early Tertiary age preserved only on its margins. Beginning about in the late Oligocene an intense episode of regional north-south extension was developed in a roughly east- west trending belt that traversed the central part of the province. This extensional belt has been termed the Mojave Extensional Belt. The extensional regime lasted from about 22-17 Ma, and is generally divided into two phases. The first, from about 22-20 Ma records intense extension via the activity of crustal-scale, simple shear, low- angle normal faults, high-angle normal faults, and extensional fracturing, along with intrusion of intermediate to silicic volcanic rocks, including episodes of explosive volcanic activity. The Peach Springs Tuff, a major regional marker unit that extends across the Mojave from Arizona to about Barstow, forms an effective stratigraphic lid on the extensional interval, (i.e., is not rotated) and is dated isotopically at about 18.5 Ma. The extensional episode was followed by an interval of high- angle normal faulting and dike emplacement from about 19 to 17 Ma, and was accompanied by at least local uplift and volcanism. Most volcaniclastic and epiclastic sedimentary units of the central Mojave block post-date the interval of extension and normal faulting, and occur as a complex of marginal alluvial and more interior lacustrine facies that formed as recently as late Miocene (ca 13 Ma). Local sequences (e.g., Lava Mountains) record deposition up to about 8 Ma, but there appears to be a regional "gap" from then until at least local successions record Pleistocene deposition. Magnetic and isotopic age studies indicate a complex history of crustal rotations, distributed across a number of discrete domains, during the past 10 Ma. Some of the youngest volcanic and geomorphic features apparently developed at "holes" formed at the edges of rotated blocks.

Jan 1, 1995

By J. -B. C. Varnier, J. A. Cremeens, D. D. Overton

"Slope failures that occurred in 2008 and 2009 at the Pikeview Quarry in El Paso County in Colorado in the United States were investigated using three-dimensional discrete element modeling methods. Large-scale block sliding failures occurred along weak bedding planes that dip steeply into an excavation located at the toe of the slope. Several mining sequences at the toe of the slope were simulated to evaluate the effects of various degrees of extraction and backfill on slope stability. Slope stability back-analysis was conducted using prefailure and post-failure conditions to determine the joint shear strength properties. Laboratory testing was also conducted to determine intact rock properties. Three-dimensional discrete element analysis was uniquely applicable to this evaluation because it allowed representation of shear stress of the basal sliding surface and afforded evaluation of several scenarios of extraction and backfill at the toe of the slope. IntroductionThe Pikeview Quarry is an aggregate quarry located approximately 8 km (5 miles) northwest of Colorado Springs in El Paso County, CO, USA. The quarry has been operated by several owners since the early 1900s. According to the U.S. Mine Safety and Health Administration (MSHA, 1986), the quarry has been a major source of construction aggregate for the Colorado Springs area for many years. In the 1950s, the site was mined by Peter Kiewit and Sons to provide limestone for the construction of the nearby Air Force Academy.A large-scale slope failure occurred on Dec. 2, 2008, which stopped mining activity at the Pikeview Quarry. A second, related slope failure occurred on Sept. 13, 2009. The second slope failure was located immediately adjacent to the initial December 2008 failure.The objective of this paper is to discuss the three-dimensional slope stability evaluation. The causal relationships of slope instability are not within the scope of this publication. However, omitting the precursors to slope failure would limit understanding of the nature of the slope stability evaluation. The regional and local geology and the history of slope instability at the Pikeview Quarry aid in understanding these precursors. Two main failure mechanisms were hypothesized during the failure investigation to explain the triggering of the 2008 and 2009 large slope failures at the Pikeview Quarry, and both of these mechanisms were given equal credence in the present study. Quarry excavation and backfilling was a key factor in the slope instability for both failure mechanisms. The excavation was backfilled in stages in an effort to provide slope support. The objective of the three-dimensional model was to evaluate the effects of extraction and backfill at the toe of the slope with respect to slope stability."

Jan 3, 2018

By C. Garcia, A. Potulska, J. J. Opazo, T. Comi

"The Sierra Gorda SCM sulfide ore processing plant began commissioning in the third quarter of 2014. Located in the Atacama Desert in northern Chile, the project is an open pit mine and Copper-Molybdenum processing plant is one of the most highly-instrumented operations in the world. It is the first hard rock operation in Chile and the second in South America to utilize HPGR technology. It is a significant Molybdenum resource especially in the first five years and the Molybdenum separation plant is complex. Designed to process up to 60 million lbs. of fine Molybdenum per year, it uses the latest technology and equipment to ensure superior recovery and product quality. It also has gas handling and alarm systems to fulfil the Company’s vision of “Zero Harm”. This paper discusses the overall operation and issues relating to the concentrator start-up. These include a review of the mineral processing design, the unified process control center and management information systems, and the correlation of geological and metallurgical performance data (mine-to-mill) for use in anticipating and preparing for treatment of diverse ore deliveries. INTRODUCTION Sierra Gorda SCM is a joint venture partnership with KGHM International (owned by KGHM SA, Poland), Sumitomo Metal Mining, and Sumitomo Corporation (Japan) operating in northern Chile. The Sierra Gorda open pit mine and Copper-Molybdenum processing plant will become one of the largest mines in the world. The need for optimum production efficiency and sustainability has driven the creation of the most modern, well-instrumented mine in Chile. This paper contains a general review of the project including geology and history. This was a very fast-track project going from greenfield site acquisition to mill start-up in 10 years. Highly-experienced and motivated engineers and technical staff have overcome many significant challenges resulting in a rapid ramp-up of production capacity. A potential expansion to double the production rate in the fourth year of operation has also been intensely studied. The Sierra Gorda property is located in the Atacama Desert of northern Chile, in the District of Sierra Gorda, Region II of Antofagasta, Chile (see Figure 1). There is almost no precipitation in this area. The mine site is generally flat with some rolling hills. The project is located about 5 km from the village of Sierra Gorda at an elevation of approximately 1,700 m above mean sea level. The village is near a highway and railway line. The city of Calama, home of the famous Chuquicamata Mine, is located about 60 km northeast of Sierra Gorda. In addition, the SG mine is in the midst of numerous other mines including Spence, Esperanza, Mantos Blancos, El Tesoro, Lomas Bayas, Gabriela Mistral, El Abra, Radomiro Tomic, and Escondida, to name a few."

Jan 1, 2015

By A. Acar, E. T. Tuzcu, Y. O. Vural

"In order to sustain the operations in chromite processing plants, due to concentrate selling cost, plants are in need of optimization in terms of capacity and recovery increase. In this study, DAMA Engineering carried out a methodological study for the purpose after a detailed sampling campaign in a plant which is situated in one the main chromite zones in Turkey; 1 The current status of the plant was investigated quantitatively with the bottlenecks, Plant overall efficiency in terms of recovery, Efficiencies of the particular equipment (mill, screen, gravity circuit), 2 A new circuit was proposed with the available equipment, 3 The new circuit that is product of modeling and experimental effort, was evaluated quantitatively as if it is a running plant, After 4-5 months continuous effort, the capacity of the plant was almost doubled (37 tph to 70 tph) and the plant overall recovery was increased from 48% to 63% by grinding adjustments and additional shaking tables theoretically. It was shown that the plant may move into profit with the minimum amount investment using readily available system. INTRODUCTION TO THE CURRENT PLANT The objective of the study is to increase the capacity and recovery of the chromite plant owned by Kurmel Group, Akmetal Inc. The complete study is given elsewhere (DAMA, 2016). The plant is situated in one of the main chromite zones in Turkey; Aladag. The current capacity of the plant is around 35 tph and the overall recovery declared by plant engineer is around 47 %. The chromite plant mainly has 3 units; crushing, grinding and gravitational circuits. The crushing circuit has 3 stages; a jaw crusher, a secondary cone crusher and a tertiary cone crusher operating in closed circuit with a vibrating screen with the aperture size of 15mm. In grinding circuit, a rod mill (2.44 6m x 3.6) and a ball mill (2.13 x 2.3m) are used for primary and secondary grinding, respectively. Derrick screens are used for the size separation. Additionally, for the size classification of the system, hydrosizers are used and each size fraction is sent to the shaking table sets. The middling product of the tables is subjected to re-grinding with a ball mill with the dimensions of 2.1 x 2.9 m. The capacity and the recovery are affected by several bottle necks; hydrosizers’ pure classification, primary/secondary mill power vs capacity problems, re-grinding product’s high slime content and so on. These combined factors made the plant to run in low recovery and capacity. Hence, DAMA was asked to make a study to investigate current status of the plant and propose a new solution to increase capacity and recovery of the plant. The conceptual flowsheet of the circuit is given in Figure 1 below."

Jan 1, 2017

By John C. Gould

GENERAL DESCRIPTION The Climax mine is located 161 km (100 miles) southwest of Denver on Fremont Pass astride the Con¬tinental Divide. Climax employs 2800 people to pro¬duce 42 670 t/d (47,000 stpd) of molybdenum ore. The Storke level, the main underground level, produces 19 970 t/d (22,000 stpd); the 600 level, the lowest ac¬tive underground area, produces 7260 t/d (8000 stpd); and the open pit presently produces 15 430 t/d (17,000 stpd), Fig. 1. A panel caving method is possible because of the favorable geology and geometry of the Climax ore body. Geology The Climax deposit is composed of three separate ore bodies formed in a massive Tertiary stockwork which penetrates Precambrian granites and schists. The ore bodies are elliptical in plan with overall dimensions of 914 x 1219 m (3000 x 4000 ft). In section the ore bodies are arcuate with an overall height of 457 m (1500 ft), although the depth is not firmly ascertained. Ore grade decreases rapidly into a highly siliceous cen¬tral core and more gradually toward the outer bounda¬ries of the deposit. The mineralization is relatively continuous in the ore zone and contains few dilution-causing waste inclusions. Molybdenite is contained in pre-ore fractures which were largely recemented during the three mineralization pulses. Postore fractures, with a spacing of a few inches to several meters (feet), render the deposit amenable to caving when it is undercut. The peripheral rock caves similarly to the ore, but the central core is competent and is difficult to cave. Caving Feasibility Successful caving action requires an undercut area of approximately 122 by 122 m (400 by 400 ft). When the undercut area exceeds this minimum distance, the arching effect is lost, and the ore column collapses. Ore extensions from the main cave area less than 122 m (400 ft) wide usually cannot be caved. Waste or low¬grade areas within the deposit can be left if they are large enough to carry weight around openings passing through them. Experience indicates that these areas should be over 30 m (100 ft) in length and width. Smaller unprofitable units of waste, such as postore dikes, are mined to avoid weight problems and result in ore dilution. Successful caving and ore recovery are dependent upon timely ore removal from the undercut zone because columns of immobile ore inhibit caving and concentrate weight onto the underlying workings. Climax experience indicates that a finger (drawpoint) spacing* averaging 10.4 m (34 ft) yields an excellent caving action and a 92.5% ore recovery at the estimated grade. An additional 15% of the ore is recovered-at a diluted grade (Fig. 2). Mine Layout The slusher drift is the primary production facility (Fig. 3). The ore is undercut and caved into the fingers,

Jan 1, 1982

"HISTORY Tube bundle systems have been used in the coal mining industry for gas detection since the 1960’s. At the time there were developments were underway for improved gas monitoring systems. The primary gas detection systems in use were the flame safety lamp, colorimetric stain tubes and the taking of samples for analysis by gas chromatography. The MSA D6 hand held methanometer was approved in the UK for statutory methane measurements in 1972 and became very popular due to its smallness and simplicity of use. In 1973 Sieger Limited also developed a BM1 methanometer which could be installed underground and relay methane concentrations to the surface. Similar developments were underway for carbon monoxide detection and one was developed in 1982 by the Yorkshire Regional Laboratory of the NCB commercial development was undertaken by Sieger Limited who named it the ‘SCO1’. The UNOR 1 carbon monoxide and methane analyzers were approved for underground use in the United Kingdom in the late 1960’s.5 The technique was developed by Dr Luft of Bergbau-Forschung Essen and the equipment was manufactured by Maihak in Germany. The apparatus was flame proof and provided an intrinsically safe output which allowed the result to be telemetered to the surface through colliery communication systems. Both analyzers work on the same principal and each of the gases adsorb a certain wavelengths of infrared radiation. The detector measures the residual transmitted radiation and determines the percentage of gas present in the sample. These flameproof versions had proven to be robust and reliable. The UNOR 1 had been used in the Saar region of Germany and Scotland in the UK. The development of tube bundles for hydraulic controls enabled the development of the tube bundle gas monitoring systems we know today. A bundle of tubes are taken underground through a shaft, a borehole or drift then distributed to the required monitoring points. This enabled the ongoing monitoring of the underground environment when the power was isolated at 1.25% CH4. In 1972 the United Kingdom (UK) National Coal Board (NCB) produced a hand book for tube bundle systems which was subsequently updated in 1974. This is a useful guideline into the operational aspects of tube bundles systems at coal mines6. Many of the processes identified are being followed today. The United State Bureau of Mines (USBM) prepared a circular similar to the NCB handbook on tube bundle systems.7 The circular was prepared by Litton C and it was targeted at the design criteria for tube bundle the systems. The circular also states that “pneumatic monitoring systems can be used to monitor for hazards that develop quite quickly”."

Jan 1, 2016

By E. McCullough

"Fostering positive community relations is a principal tenet of sustainable development, but conflict can test even the strongest of relationships. While conflict surrounding resource development projects has the potential to be destructive, it may actually pose opportunities for partnership and collaboration when managed effectively. It is therefore not only important for mineral resource professionals to recognize the financial and social costs associated with legal disputes among project stakeholders, but also for them to prioritize resolving social conflict using alternative channels that allow for win-win outcomes. A variety of dispute prevention and resolution techniques do exist, but successful practice is not yet the norm. Here we review the sources and consequences of social conflict in the context of mining projects, and discuss typical and alternative approaches to conflict management in light of their respective benefits and drawbacks. We also suggest implementation of conflict resolution lessons into resource-related university curricula as a critical step in the continued shift of the mining industry towards sustainable development. INTRODUCTION Social conflict is a recurring phenomenon in the human experience and serves as a critical mechanism for change [1-5]. It occurs when two or more parties perceive that their needs and interests are in direct competition with each other and believe they cannot be reasonably accommodated. Conflict over mineral resource activities, particularly the distribution of benefit and risk, has become an omnipresent issue throughout the world [6-7]. Contrary to common perception, however, conflict can be a positive, transformative encounter in some instances [2, 8-11]. For this to be possible, society and its acting agents must reconsider how to address conflict. What if conflict were only a challenge to try something different? It is well-understood that the mineral industries must manage an increasingly significant volume of social conflict [12-13], which can be a costly externality impacting all of whom are involved. We commonly consider conflict in a two-party sense, but this deeply over-simplifies the true number of stakeholders. Besides mining companies and community members, government policymakers and regulators, public interest groups, and judicial systems, to name a few, bear the social expenses of poorly-managed conflict [4, 6, 12, 14-22]. It is desirable for interested parties to participate actively in a constructive way to pursue their needs rather than contributing to adversarial relationships, operational problems, and legal costs [1, 4, 12]. Some mining companies are of course already working to develop more productive relationships with their stakeholders, but there is undoubtedly still a gap between ideal and common practices. This gap can inhibit collaboration and threaten the potential for mutually-beneficial resolutions to complex social issues [12]."

Jan 1, 2016

By E. McCullough

"Fostering positive community relations is a principal tenet of sustainable development, but conflict can test even the strongest of relationships. While conflict surrounding resource development projects has the potential to be destructive, it may actually pose opportunities for partnership and collaboration when managed effectively. It is therefore not only important for mineral resource professionals to recognize the financial and social costs associated with legal disputes among project stakeholders, but also for them to prioritize resolving social conflict using alternative channels that allow for win-win outcomes. A variety of dispute prevention and resolution techniques do exist, but successful practice is not yet the norm. Here we review the sources and consequences of social conflict in the context of mining projects, and discuss typical and alternative approaches to conflict management in light of their respective benefits and drawbacks. We also suggest implementation of conflict resolution lessons into resource-related university curricula as a critical step in the continued shift of the mining industry towards sustainable development. INTRODUCTION Social conflict is a recurring phenomenon in the human experience and serves as a critical mechanism for change [1-5]. It occurs when two or more parties perceive that their needs and interests are in direct competition with each other and believe they cannot be reasonably accommodated. Conflict over mineral resource activities, particularly the distribution of benefit and risk, has become an omnipresent issue throughout the world [6-7]. Contrary to common perception, however, conflict can be a positive, transformative encounter in some instances [2, 8-11]. For this to be possible, society and its acting agents must reconsider how to address conflict. What if conflict were only a challenge to try something different? It is well-understood that the mineral industries must manage an increasingly significant volume of social conflict [12-13], which can be a costly externality impacting all of whom are involved. Conflict is commonly considered in a two-party sense, but this can deeply over-simplify the true number of stakeholders. Besides mining companies and community members, government policymakers and regulators, public interest groups, and judicial systems, to name a few, bear the social expenses of poorly-managed conflict [4, 6, 12, 14-22]. It is desirable for interested parties to participate actively in a constructive way to pursue their needs rather than contributing to adversarial relationships, operational problems, and legal costs [1, 4, 12]. Some mining companies are of course already working to develop more productive relationships with their stakeholders, but there is undoubtedly still a gap between ideal and common practices. This gap can inhibit collaboration and threaten the potential for mutually-beneficial resolutions to complex social issues [12]."

Jan 1, 2016

By T. P. Lim, J. A. Cherry, A. J. Vivyurka, R. D. Blair

The first phase of an investigation of the chemical composition of subsurface water and the hydrochemical processes in and near the abandoned Nordic tailings at Elliot Lake, Ontario, was conducted in 1979. The Nordic area, which covers 85 hectares with an average thickness of 10 m and overlies deposits of permeable glaciofluvial sand, contains 15 million tonnes of tailings with 4-7% pyrite. Nests of standpipe piezometers installed at three locations on the tailings indicate that seepage through the tailings is predominantly downward. The water table in the tailings slopes from the north side of the tailings dam, where it is within a metre of surface to the south side near the tailings dam, where it is approximately 10 m below surface. When they were deposited during the period from 1957 to 1968, the tailings contained pore water at a pH of about 8 resulting from neutralization treatment in the mill. At the three piezometer nests, the pH increases from 3 near the water table to 7 near the bottom of the tailings. The pore water in the tailings is gradually becoming acidic because of oxidation of pyrite in the zone above the water table. The concentrations of Fe, Mn, Ni, Co, Zn and Pb are high in the shallow acidic zone and much lower in the deeper neutral zone. Concentrations of 226Ra range from 30 to 230 pCi/L. At the three sites concentrations of major ions in the permeable sand a short interval beneath the tailings indicate 226Ra presence of tailings seepage, but Ra concentrations are only 2 to 8 pCi/L. Greater downward penetration of 226Ra probably occurs in the area closer to the tailings dam. Monitoring of a network of multilevel bundle piezometers installed in permeable sand that extends southward from beneath the tailings indicates the occurrence of a plume of tailingsderived water containing high concentrations of Ca, Mg, Fe and S04 that extend 400 m downgradient from the tailings dam. The total dissolved solids decline from 20,000 at the dam to less than 500 mg/L at the identifiable downgradient periphery of the plume. A much smaller but distinct plume of water with high heavy metal concentrations and 226Ra in the range of 10 to 130 pCi/L extends from beneath the dam for a distance of 15 m. The pH of the groundwater in this heavy metal-radium plume is 4 to 5, whereas elsewhere in zone of groundwater with high dissolved solids, the pH is above 6. The results of this investigat, suggest that the heavy metals an 226 Ra are relatively immobile in the perme-

Jan 1, 1980

By Richard B. Muter

1. Coal Preparation Terms (General) 1.1 Coal preparation-collectively, physical and mechanical processes applied to coal to make it suitable for a particular use. 1.2 Run of mine (ROM coal)-coal produced by mining operations, before screening, crushing, or preparation. 1.3 Raw coal-coal that has received no preparation other than possibly screening or crushing. 1.4 Raw coal feed-raw coal supplied to a plant or machine, in which it undergoes some form of preparation. 1.5 Coal cleaning-the treatment of coal to lower the mineral matter content (ash). 1.6 Cleaned coal; clean coal-coal produced by a cleaning process (wet or dry). 1.7 Middlings-a product of coal preparation that, by reason of its ash content, is intermediate between coal and discard. It follows therefore that the relative density of middlings is intermediate between those of coal and discard. Middlings may be reprocessed. 1.8 True middlings; bone comparatively high ash particles so nearly homogeneous that their quality cannot readily be improved by crushing and recleaning. 1.9 False middlings; interbanded middlings-middliigs in which the particles consist of bands of coal and shale, and from which the coal may be liberated by crushing 1.10 Refuse-waste material in the raw coal which the cleaning process removes. (See also discard; dirt; stone, No. 1.12). 1.1 1 Reject-the material extracted from the feed during cleaning, for retreatment or discard. 1.12 Discard; dirt; stone--the material extracted from the raw coal and finally discarded. 1.13 Recirculation-the operation in which the whole or part of a product from a process is returned to the feed to the process, e.g., the return of the crushed overflow from a screen to the screen feed for rescreening. 1.14 Foreign coal-coal received at a preparation plant from a colliery other than that to which the plant is attached. 1.15 Imported coal-coal coming from a foreign country, or other state within the country. 1.16 Low-grade coal combustible material that has only limited uses owing to undesirable characteristics (eg., ash content or size). 1.17 Segregation-partial separation of a previously mixed material into its constituents, occurring as a result of differences in either particle size or relative density. 2. Cleaning Characteristics 2.1 Washability-amenability of coal to improvement in quality by cleaning. 2.2 Float-and-sink analysis-the division of a sample into relative density fractions with defined limits, the proportions of the fractions being expressed as percentages of the total sample, commonly with an indication of the ash percentage (and other characteristics, if required) of each fraction.

Jan 1, 1991

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 D. Burkhard, E. Herth

"Geologic structure and properties are the primary influences on ground control at San Juan Mine. Geotechnical properties of the coal seam and near seam strata, the presence of above seam sandstone paleochannels and water, are thought to be the leading contributors to ground control issues. It is the confluence of specific geologic conditions rather than isolated features that most often produce adverse conditions. Altering primary support installation and mining practices are done when unfavorable conditions appear during development. Installation of standing support is most often employed when adverse conditions develop at later dates. REGIONAL GEOLOGIC SETTING San Juan Mine is an underground coal mine located on the western flank of the San Juan Basin in northwest New Mexico. The San Juan Basin is contained within the south eastern region of the Colorado Plateau geographic province. The basin contains multiple coal bearing formations and is also a major producer of coalbed methane gas. San Juan Mine is stratigraphically located within the Fruitland Formation. From a prominent monocline marking the western boundary of the basin, the formation dips gently to the east towards the center of the basin. The primary driver for the deposition of coal bearing formations within the basin was the Cretaceous Inland Seaway. Multiple cycles of transgression and regression driven by sea level rise and tectonic subsidence deposited rock types from marine shale’s to terrestrial sandstones. The Sevier Mountains to the west of the seaway likely provided the sediment supply for the Fruitland Formation. During a major cycle of seaward regression, the Pictured Cliffs Sandstone Formation prograded over the marine Lewis Shale. The coal bearing Fruitland Formation followed, being deposited over the Pictured Cliffs. The Fruitland Formation contains numerous coal seams and is approximately 400 ft. thick at San Juan Mine. Fruitland Formation coal seams are numbered in order of superposition. San Juan mines the Fruitland number 8 seam. Dating has placed the age of the seam at approximately 73 to 75 million years [1]. LOCAL GEOLOGIC SETTING The Fruitland 8 coal seam dips approximately 1-3 degrees towards the E-NE. At San Juan Mine, the seam displays a mild plunging, synclinal structure, the axis roughly bisecting the mine from west to east. The number 8 seam consists of five coal benches and four volcanic ash band partings labeled sequentially based on superposition. The lowest portion of the seam is the 8-1 bench, the 8-2 through 8-5 benches following sequentially to the top of the minable seam. Ash band partings are labelled P1 through P4 from bottom to top respectively. The presence of distinctive partings and benches allows for seam definition and aids in identifying mining horizons."

Jan 1, 2016

By Cahit Helvaci, Robert B. Kistler

Borates are defined as &apos; &apos;salts or esters of boric acid; a compound containing the radical B203" (Bates and Jackson, 1987). Borates are defined by industry as any compound that contains or supplies boric oxide. A large number of minerals contain boric oxide, but the three that are the most important from a worldwide commercial standpoint are: borax, ulexite, and colemanite. These are produced in a limited number of countries (Fig. 1), dominated by the United States and Turkey, which together furnish about 90% of the world&apos;s borate supplies (Lyday, 1991). Production in the United States originates in the Mojave Desert of California; borax and kernite are mined by US Borax from its large deposit at Boron. Borate containing brines are pumped from Searles Lake by North American Chemical Co., and a limited amount of colemanite is mined by Newport Mineral Ventures from Death Valley. Turkish production is controlled by Etibank, the national mining enterprise, which supplies most of the commercially traded ulexite and colemanite from mines in the Bigadic and Emet Districts, plus borax from the huge deposit at Kirka. Borate minerals have been employed in a wide range of uses for many centuries, dating from at least the 8th century when they were used primarily as a flux for assaying and refining gold and silver. Their valuable properties and relative rarity soon stimulated international trade in borates; Marco Polo claimed to have trans- ported Chinese borate minerals from Tibet to Europe, and his home port of Venice was the center for borate imports and use throughout the Middle Ages (Travis and Cocks, 1984). Borates were traded at relatively high prices for highly specialized applications into the late years of the 19th century. At that time they were being used for medicines, food preservatives, ceramic glazes, and in expanded applications as metal fluxes. It was in that era that borax was discovered in Clear Lake, CA, and more significantly, the playa deposits of western Nevada were found and exploited by F.M. "Borax" Smith and others. Research and development by Smith&apos;s company began the evolution of borates from a high priced commodity with a few specialized uses to the relatively low priced chemical raw material used in large tonnages on a worldwide scale today. Borates are often defined and sold by their boric oxide or B,03 content, and most statistical data are listed in tons of B203. Of the products sold by the industry, borax pentahydrate or 5 mol and boric acid are the most commonly traded commodities. Boric acid plants are operated by all of the major borate producers. Glass fiber insulation is the major end use in the United States followed by textile glass fiber and borosilicate glass, detergents, and ceramics. Detergent usage continues to be a major end use in Europe.

Jan 1, 1994