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By L. Yuan, A. C. Smith

MFIRE, the mine-fire simulation program of the U.S. National Institute for Occupational Safety and Health (NIOSH), is widely accepted as a standard for assessing and predicting the impact of a fire on the mine ventilation system and the spread of fire contaminants in coal and metal/nonmetal mines. MFIRE has been used by U.S. and international companies to simulate fires for planning and response purposes. It is a dynamic, transient-state, mine ventilation network simulation program that performs normal planning calculations. It can also be used to analyze ventilation networks under thermal and mechanical influence, such as changes in ventilation parameters; external influences, such as changes in temperature; and internal influences, such as a fire. The program output can be used to analyze the effects of these influences on the ventilation system. Since its original development by Michigan Technological University for the former U.S. Bureau of Mines in the 1970s, several updates have been released. In 2012, NIOSH completed a major redesign and restructuring of the program with the release of MFIRE 3.0. MFIRE’s outdated Fortran programming language was replaced with an object-oriented C++ language and packaged into a dynamic-link library. However, the MFIRE 3.0 release included no change or improvement to the fire-modeling algorithms inherited from its previous version, MFIRE 2.20. This paper reports on improvements that have been made to the fire-modeling capabilities of MFIRE 3.0 since its release. These improvements include the addition of the t-squared fire as a new fire-source model and the ability to include available heat-release-rate versus time data, the addition of a moving fire source for conveyor-belt fire simulations, improvement of the fire location algorithm, and the identification and prediction of smoke-rollback phenomena. All of the improvements discussed in this paper will be included in MFIRE 3.1 and released by NIOSH in the near future.

Jun 1, 2016

By Dan Nilsson

The Luossavaara "Research Mine" is the First large, offensive effort in mining research conducted by the Swedish state. Over a 5-year period, 1981-1985, the Swedish Mining Research Foundation works in the "Research Mine" on the development of mining systems and equipment. The purpose is to bolster the competitive strength of the Swedish mining and mining equipment industry. The previosly abandoned Luossavaara mine in Kiruna, in northern Sweden, has been reopened especially For the project. Ore is mined and sold during the course of the research project. The income from the sales covers a large portion of the costs for teh research.

Jan 1, 1985

By J. J. Harrington

Manufacture and mixing of flotation reagents at the mill site provides an economic and convenient method of maintaining a continuous and fresh supply of reagents for concentrator operations. Kennecott Copper Corporation has designed and built a Reagent Mixing Facility which incorporates new ideas in construction and fail-safe operation at its Nevada Mines Division concentrator. Highly sophisticated automation and instrumentation techniques are combined with safety procedures and equipment that meet and exceed all Federal and State safety requirements. Installation of the facility has minimized hazardous working conditions for the operator, isolated reagent preparation from production personnel and equipment, and reduced the time required for reagent preparation from three shifts to one shift per day.

Jan 1, 1976

By Pedro Pablo Vasquez Coronado, Victor Octavio Tenorio

"Productivity in openpit mines depends on the observation and adjustment of critical phases of operation. As one of the most expensive activities in openpit production, optimization of the load-haul-dump cycle becomes a critical aspect to address. Using simulation techniques and an ad hoc methodology for analyzing the key performance indicators (KPIs) involved in the production cycle, it is possible to compare actual performance and ideal operating conditions and develop an alert system when operating conditions are subpar with the expected production standards.Study objectivesThe purpose of this work is to implement a methodology for optimizing the haulage cycle by identifying and adjusting the factors that affect the cycle time performance. The methodology includes the implementation of an alert system emitting warning messages when one or several trucks are idle, or when the time of completing production tasks is falling outside a predefined value. This alert is compared with predefined KPIs and is used to determine corrective actions.The alert will:• Demonstrate the benefits of time analysis and how this affects the productivity results in the mining process.• Demonstrate the utility of KPIs in the extraction of ore and waste and how they help to identify equipment performance.• Develop software interfaces to visualize the alert system in order to increase awareness in the operations supervisor.• Make recommendations for future applications that can help improve the productive process of an openpit mine.Problem statementTransporting material in openpit mines may require the utilization of high capacity equipment, which also represents an elevated cost per hour. A supervisory system that controls the adequate distribution of trucks and shovels by providing the operational status of each unit shall contribute to increase productivity while subsequently reducing operational costs. Nevertheless, there is an issue with certain dead times such as queue trucks for loading, trucks waiting to be loaded and other delays. This creates inefficiencies in fuel consumption and energy utilization.For the identification of delays, a contextbased system could be required in order to help optimize the haulage cycle time by controlling delays in loading, haulage and dumping. After applying this methodology, efforts were focused in correcting out-of-standard performance values."

Jan 1, 2015

By R. K. Mehta

Stirred ball mill feed used for ultra-fine grinding application requires fine size separation step in order to avoid over-grinding. This paper presents the selectivity curves obtained for oil shale and complex sulfide are using 460 mesh (30 micron) Derrick vibrating screen. Interestingly, the selectivity curves generated from the experimental data were monotonic increasing selectivity for oil shale but not monotonic increasing for sulfide ore. These non-monotonic increasing selectivity curves were similar to the mechanical air separator fishhook selectivity curves described by Luckie and Austin. A 2 stage classification in series model describing the fishhook selectivity curve has been used in proposing the classification mechanism for fine vibrating screens when heterogeneous multicomponent sulfide are undergoes size separation.

Jan 1, 1992

By M. S. Celik

Salt-type minerals are usually floated with anionic and cationic collectors. Flotation of a number of systems has been reported to remarkably increase above the concentrations where surface precipitation of collector-metal salt is initiated. Some studies attribute this phenomenon to heterocoagulation of oppositely charged collector colloid and mineral particles. In this study, examples are given on boron minerals to demonstrate that the high flotation region occurs when particles, colloidal precipitates and bubbles acquire a similar charge which induces Van der Waals interactions resulting in hydrophobic bonding.

Jan 1, 1995

By Donald E. Steif

The Tower Mill is a vertical stirred mill using steel balls or pebble grinding media. This device is designed for fine grinding at significant power savings when compared to conventional tumbling mills. This paper describes the machine, its operation, application, and performance for wet grinding.

Jan 1, 1984

By J. J. Carlson

An important factor in filtration is the zeta potential of the particles. Since zeta potential is a function of pH, it was hypothesized that adding CO would lower the pH and zeta potential. A reduction in zeta potential should enhance the flocculation and increase the filtration rate. This paper describes the effects CO on the zeta potential of both pure hematite and samples of hematite concentrate from an iron ore processing plant. A significant shift in the zeta potential, as a function of pH, curve was observed. These results point to an alteration of the particles mineralogy, not just a shift in pH.

Jan 1, 2010

By Dapeng Zhang, Brij Moudgil, Shibo Zheng, Chunhui Feng, Wenyi Song, Wending Xiao, Patrick Zhang

The separation of dolomite from phosphate is the most challenging problem in phosphate mineral processing. More than 50 percent of the phosphate reserve in Florida contains too much dolomite to process using current industry practice. The current way of producing phosphoric acid requires less than one percent MgO in the phosphate concentrate as the feed material. The Florida Industrial and Phosphate Research (FIPR) Institute has collaborated with worldwide experts in the field to address this issue. As a result, the industry is now offered three feasible options. Option I entails three methods for reducing MgO content in the concentrate from the Crago process, including adding a dolomite depressant in the rougher flotation step, dolomite flotation of the cleaner concentrate, and scrubbing the cleaner concentrate in quartz sand, which could reduce MgO content in the final concentrate by 20 to 40 percent. Option II involves the crushing and grinding of high-dolomite phosphate pebbles followed by dolomite flotation at slightly acidic pH using a new collector that does not require phosphoric acid as a phosphate depressant, achieving a final concentrate analyzing less than 0.9 percent MgO at about 87 percent P2O5 recovery. Option III is a gravity separation technique using an innovative separation jig, and its full potential remains to be demonstrated.

By Paul Piercy

Bougainville Copper Limited operates a porphyry copper mine in Papua New Guinea. The waste dumps will contain 400 to 500 million tomes of rock at the end of nine life. The paper describes investigations carried out over a period of en years and highlights the problems associated with waste lump leaching in a tropical and seismic active region. Aspects discussed include mineralogy, chemical reaction, leaching rates, microbiological activity, dump stability, dump structures, effects of tropical conditions, seismic activity and topography.

Jan 1, 1981

It was shown previously that chalcopyrite could be floated without using a collector when the ore samples were treated with small amounts of sodium sulfide (Na2S). The role of this reducing agent was considered to be one of a cleaning agent to remove the hydrophilic surface oxidation products. It was later found, however, that the collectorless flotation could be obtained only under slightly oxidizing conditions. This might suggest that collectorless flotation is not due to the clean surface but to the elemental sulfur which may be present as an incipient oxidation product. Therefore, a search for the presence of elemental sulfur has been made by means of ESCA, mass spectrometry, and solvent extraction followed by W spectrophotometric analysis. On the basis of these results, the mechanisms of collectorless flotation are discussed.

Jan 1, 1983

By David C. Sanders, Rodney H. Sergent, Victor C. Storhok, Batric Pesic

In the past, considerable work has been done on the applica­bility of bromine chemistry in the extractive metallurgy of gold. While the research in this area spanned from the understanding of fundamental dissolution chemistry to the demonstration on a pilot plant scale, no work was done on the fate of discarded leach residues with respect to the stability of accompanied heavy met­als. Because the leach residues were the discharge products of halide/halogen systems, the general perception was that the con­tained heavy metals would be more susceptible to dissolution than in typical sulfate systems, thus causing increased environ­mental problems. A long-term study was performed to obtain the information on the stability of heavy metals contained in the leaching residues after bromination. The study involved the leach residues of three different ores, as well the precipitates prepared from the synthetic solutions (not presented here). The conclusion of this study is that the stability of solid residues produced by bromination leaching was same as the sta­bility of leach residues produced by leaching in the sulfate sys­tems. Also, contrary to the past expectations, the presence of bromide ions in the test solutions (up to 1 g/L NaBr) did not enhance the solubility of heavy metals (Cu, Zn, Pb, As) beyond the solubilities found in the sulfate systems.

Jan 1, 1998

By Yoginder P. Chugh

The mining methods used at many opencast brown coal mines in Poland and the German Democratic Republic are not used in the US to any extent. Using bucketwheel excavators, bucket chain excavators, and mobile conveyor bridges, the mining techniques may offer efficient, economic alternatives for operators in the US. These mining methods were observed during visits to several open-cast brown coal mines in Eastern Europe over the past two years. Some of them may be particularly attractive for the large lignite mines being planned in North Dakota and Texas.

Jan 11, 1980

By X. F. Wang, K. B. Fu

"We investigated experimentally and theoretically the adhesion behavior of Acidithiobacillus ferrooxidans (At.f6) to different copper sulfides. The adsorption experiment showed cell densities of 10.74 × 106, 8.06 × 106, 3 × 106, 4.24 × 106 and 2.35 × 106 cells/cm2, respectively, on the surfaces of djurleite, bornite, covellite and chalcopyrite. The order was: djurleite > bornite > covellite > chalcopyrite. The results indicate that A. ferrooxidans has high affinity to adhering to the surfaces of djurleite and bornite in the acidic region. But the thermodynamic approach predicts that there will be no attachment of At.f6 to the surfaces of bornite, covellite and chalcopyrite. This discrepancy is due to the inadequate description of electrostatic interactions. The DLVO approach shows attachment of At.f6 to the surfaces of djurleite, bornite and covellite, and repulsion between At.f6 and chalcopyrite. The DLVO approach can explain the differences of the adhesion of At.f6 to different copper sulfides.IntroductionThe bioleaching of low-grade copper sulfides is a developing technology, and has been applied successfully to the extraction of copper from secondary sulfide minerals (Watling, 2006; Pradhan et al., 2008; Akcil and Deveci, 2010). The bioleaching of chalcopyrite is still a major challenge (Bosecker, 1997; Brierley and Brierley, 2001; Ehrlich, 2004). A study conducted at Billiton Process Research showed the following preferential order of sulfide minerals leached by mesophilic cultures: chalcocite > bornite > cubanite > covellite > pyrite > enargite > carrolite >> chalcopyrite (Dew et al., 1999). The bacteria can solubilize the valuable metals from the orebodies by the oxidation of minerals. Bacterial adhesion plays a critical role in the bioleaching processes (Chen et al., 2008; Harneit et al., 2006). Generally, bacterial adhesion can be explained by surface thermodynamics and the extended DLVO theory, in which the adhesion energy between the bacteria and mineral is calculated as a function of the separation distance (Sharma, 1999; Bos et al., 1999; Sharma and Hanumantha Rao, 2002). These methods take into account Lifshitz-van der Waals interactions, electrostatic interactions and acid-base interactions (Busscher et al., 2010). These interactions are very well understood and formulated in mathematical equations. The most important inputs in these calculations are surface energy, surface charge, electron donation and electron acceptance on the bacterial cell surface and mineral."

Jan 1, 2015

By Jessica M. Wempen, Dallan J. Coons, Bailey S. Simmons, Michael K. McCarter

Gate-road yield pillars are subject to increases in stress as the longwall face advances toward the pillars. Increases in pillar stress can be inferred by measuring roof/floor convergence surrounding the pillar. This paper presents a case study of convergence measurements surrounding a two-entry gate-road yield pillar in the head gate of a coal mine. Convergence was measured over a six-month period, as the longwall face approached and passed the pillar. Measurements were collected at eight locations surrounding the pillar: four locations in the travelway and four locations in the beltway. In addition to convergence stations, the pillar was instrumented with an array of seismometers. In future analysis, convergence measurements will be compared to seismic interpretations. This study reports on the geology and geometry of the longwall panel, the roof and rib support, the change in convergence over time, and the physical properties of coal and parting. Additionally, the convergence measurements are compared to pillar safety factor distributions derived from a finite element model. The convergence measurements contribute to the knowledge of when yield pillar abutment loading, caused by the encroaching longwall, is first detected. Knowing when abutment loading begins to yield the pillars can help define when standing secondary roof supports are required for head gate entry support.

By R. S. Rickard, B. Ball

INTRODUCTION Pyrite responds to the flotation reagents normally used in sulfide flotation systems. For example, it is well known that xanthates are good collectors for pyrite, and cyanide, sulfide and hydroxides are depressants. The particular response of pyrite to these and other reagents is used to advantage in several areas. By careful control of depressant and collector additions, pyrite can be separated from other sulfides by flotation. This is of obvious importance in base metal sulfide flotation where pyrite depression is often required to produce a marketable grade concentrate. Flotation of pyrite has long been employed in gold ore processing, and it appears that pyrite flotation from coal may become important in the future (1). It is probably a direct consequence of this relative ease of separation, plus the low economic value of pyrite, that detailed investigations on the chemistry of its flotation and depression are relatively limited. Gaudin (2,3,4) and others, in the mid-1950's published data on the flotation of pyrite, but detailed or consistent mechanisms were lacking. In 1955, Sutherland and Wark's book (5) was published, in which a large amount of empirical information, especially earlier work of Wark and Cox, was presented. More recently, in 1968, significant contributions have been made by Fuerstenau et al. (6,7) and Majima and Takeda (El. These papers introduced electrochemical explanations of the effects observed with xanthate and cyanide ions.

Jan 1, 1976

By J. O. Marsden, M. M. Botz

"Heap leaching is responsible for approximately 21 percent — 3.9 million metric tonnes — of copper production and 9 percent — 270 metric tonnes or 8.7 million ounces — of gold production worldwide. Given metal price assumptions of $2.25/lb for copper and $1,250/oz for gold, these portions of global production generate revenues of $19 billion and $11 billion, respectively. For heap leaching operations, small changes in metal extraction efficiency drive substantial changes in operating cash flow margins and affect project viability. Variances in actual metal production from budgetary forecast estimates can mean the difference between success (value creation) and failure (value destruction) for a heap leach project. Despite the importance of heap leaching for copper, gold and silver production worldwide, there are no clear standards or guidelines for developing heap leach metal production forecasts. There is considerable variation among copper and gold producers that apply heap leaching technology as to how such modeling and forecasting is applied at the operational level. There are a variety of different modeling approaches that can be utilized for this purpose. However, these vary in cost, complexity, time to implement, accuracy and credibility. This paper explores the options available for heap leach modeling, discusses the advantages and disadvantages of each approach, and provides guidelines/recommendations for an effective approach to modeling metal production at existing or planned heap leach operations. The paper also provides expected ranges of variation between actual and modeled production.Introduction and backgroundHeap leaching is fundamentally different from other methods used in the processing and extraction of gold, silver, and copper and other base metals. Considered globally, the technique represents a relatively modest portion of production compared with the primary technologies applied for copper and gold recovery. Approximately 21 percent of copper production is attributable to heap leaching techniques, as opposed to 79 percent by conventional crushing, milling, flotation, smelting and refining. For gold, about 9 percent of global production is obtained from heap leaching, compared with over 90 percent by various combinations of crushing, milling, gravity concentration, flotation and/or agitated cyanide leaching processes or, in the case of placer deposits, by direct gravity concentration and/or amalgamation. In cases where heap leaching is the primary source of metal production, the ability to accurately forecast metal production may be the difference between commercial success and failure for a project or operation. Heap leaching — of crushed or run-of-mine ore — is also commonly applied as a secondary process to treat lower-grade ores from a given deposit, such as at Chuquicamata in Chile, Cerro Verde in Peru and Bagdad in Arizona for copper extraction and at Twin Creeks and Carlin, both in Nevada, for gold. In such cases, heap leaching provides incremental revenue to supplement the primary process. Variances between actual and forecast metal production from heap leaching as a secondary process are typically tolerated and absorbed within the overall site metallurgical balance. In either case, accurate production modeling and forecasting creates significant value for heap leaching projects, is essential for managing production expectations — timing and volumes — and is likely to be a critical success factor for most heap leach projects."

Jan 1, 2017

By M. Jaramillo

The Expanded West Pit of the 15 MTY El Cerrejon Mine will reach a maximum depth of270 m. Slope stability evaluations indicated that this depth, combined with structural complexities, would create a serious risk of footwall instability. Stabilization call for anchoring the footwall by installation of cable bolts, 9 to 33 m long, of multi-strand steel cable. For each of the 15 m high mining benches, a row of bolts is installed with variable horizontal spacing, depending on dip magnitude. Cable bolting is now standard operating practice and 11800 bolts will have been installed during the life of the mine.

Jan 1, 1995

By Kugan Kuganathan

Pastefill is being increasingly considered as a viable fill by underground mine operators. Mount Isa Mines (MIM) chose pastefill to fill their deep copper mine (>1200m deep) after considering various other types of fill in the feasibility studies. During the development of pastefill at Mount Isa (1996 - to date) the Mining research of MIM developed two cost effective, reliable and simple testing methods and used. With the help of these tests the behaviour of MIM pastefill was studied thoroughly and the most optimum mix design for pastefill was achieved. The MIM's pastefill plant, the largest in the world with a capacity of 450rph continuous pastefill production and unique in the process used, has been built (1998-1999) to produce pastefill to the specification developed through extensive testing program using the Flow Cone and Mini Slump cone Tests. As a consequence, it is now possible to select the optimum pastefill mix designs and predict their flow behaviour in bore holes or pipe lines using only these tests. MIM Blended Tailings Pastefill (BTPF) uses 3-3.55 Portland cement and the cost of pastefill is much lower than the total tailings pastefill, that could be produced from the same tailings.

Jan 1, 2001

By M. A. Georgiana

The Emerald Mine No. 1, located in southwestern Pennsylvania is presently owned and operated by Cyprus Minerals of Englewood, Colorado. High volatile bituminous coal is produced from the Pittsburgh No. 8 seam by both longwall and continuous mining methods. The coal is then mechanically cleaned on site utilizing a combination of heavy media separation and froth flotation methods prior to shipment to both the metallurgical and steam coal markets. The Emerald Mine is presently producing 2.8 million raw tons per year, while liberating approximately 6.5 million ft per day of methane. In light of the methane liberation rate, ventilation requirements have been significant and have always received utmost priority and consideration. Figure 1 shows the locations of the current and proposed shafts. Emerald utilizes two Joy (Model M120-58-880) axial vane fans on shafts No. 1, No. 2R, No. 3R. (one operational, one stand-by) The use of computer simulation of various ventilation networks was initially employed during the design stages (1975-1977) and has been utilized through the present. Obviously, the initial planning, under the ownership of Lykes Resources, was strictly theoretical. All the quantities were proposed and the resistances were based on prior simulations of other mines and mathematical formulas. However, once the mine was moderately developed, actual values were recorded and utilized. This procedure consists of: a) layout pressure and quantity points b) measure all air quantities where an air course splits

Jan 1, 1986