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By Isa Nozari, Asghar Azizi

The extraction of copper from sulfate leaching liquor was investigated using Acorga M5640 commercial extractant in kerosene diluent. The effects of important factors such as the extractant concentration, contact time, liquor pH, and temperature were examined on the separation factor Cu/Fe and the extraction efficiency of copper. The results indicated that nearly 92.32% copper was extracted from the leach liquor at 10% (v/v) Acorga M5640 concentration, 1:1 aqueous/organic phase ratio, 20 min contact time, pH 2.5, and 25 °C temperature. Also, the further increase in the values determined had no positive impact on the extraction of copper. Under these conditions, the separation factor Cu/Fe was obtained about 29.64. The extraction mechanism was studied applying the plot of log(D) – 2 pH against log [extractant] and it proved the association of two moles of Acorga M5640 for the extraction of one mole of copper. McCabe–Thiele diagram was constructed and a single-step extraction process was found to obtain themaximumextraction. Additionally, thermodynamic datawere investigated and the extraction process was endothermic and non-spontaneous at 298 K (ΔH = 27.05 and ΔG = 15.52 kJ mol−1). Moreover, the loaded organic phase was stripped and 80.94% copper was recovered after 20 min using 10% sulfuric acid concentration.

Jul 31, 2020

Small-scale flotation tests have been conducted to examine the effect of froth stability on the column flotation of finely pulverized coal. The results of this study demonstrate that the upgrading of coal in a column flotation cell can be significantly improved when froth stability is properly controlled through the manipulation of appropriate operating variables. Several important correlations have been established from the test data. These relationships should prove useful for the design and ore ration of column flotation cells for fine coal processing.

Jan 1, 1991

By K. J. Ma, L. Ma, J. Stankus

The paper’s intent is to evaluate pumpable foamed cementitious mine seal material by determining the shear strength of a generic foamed cement in the laboratory. The tests were conducted on three scales: small-size specimens, consisting of cylinders 4 in. in diameter or 3-in. prisms; double shear medium- size specimens, measuring 13 in. by 10 in.; and large-size specimens, measuring 6 ft by 4 ft by 4 ft. The small-specimen shear tests resulted in average peak and residual shear strengths of 128 psi and 56 psi, respectively. These values are fairly consistent with published shear strengths of foamed cementitious material. However, the other two types of shear tests did not achieve the shear strength of a typical pumpable foamed cementitous mine seal material. Based on the test results, the paper concluded that the structural analysis of cement foam seals should be based on their residual shear strengths, approximately 20 psi, determined from small laboratory-scale tests of the cement foam/concrete or rock interface, and not be based on their peak strengths.

Jan 1, 2015

By F. Freme

Coal production in the United States in 2009 de-creased to 972.8 Mt (1,072.8 million st), according to preliminary data from the U.S. Energy Information Administration (EIA). This is a decline of 8.5 percent, or 90 Mt (99.1 million st) below the 2008 record level of 1.06 Gt (1,171.8 million st) (Table 1). In 2009, U.S. coal consumption decreased in all sectors while total coal stocks increased for the year. Coal consumption in the electric power sector in 2009 was lower by 10 percent, while coking coal consumption decreased by 30.6 percent and the other industrial sector declined by 16.6 percent. The commercial and institutional sector (which prior to 2008 had been called ?residential and commercial?) is the smallest of all the coal-consuming sectors and it declined by 8.4 percent in 2009. (Note: all percentage change calculations are done at the short tons level.) U.S. coal exports fell from the 2008 levels, while coal imports decreased for a second year in a row. The decline in coal consumption during the year was the consequence of the domestic economic conditions combined with the weather in 2009, which resulted in lower demand for electricity. Preliminary data show that total generation in the electric power sector (electric utilities and independent power producers, including useful thermal output) in the U.S. decreased in 2009. Coal-based generation also decreased, resulting in a 94-Mt (104-million st) drop in coal consumed in the electric power sector. Coal use in the non-electricity sector decreased by 20.1 percent to a level of 57.9 Mt (63.9 million st). The international coal markets for most of 2009 were reflective of a worldwide economic slowdown. U.S. coal exports and imports both declined for the year. U.S. coal exports totaled 53.6 Mt (59.1 million st). This is a decrease of 20.3 Mt (22.4 million st) from the 2008 level.

Jun 1, 2010

By J. F. T. Agapito

Stability in Exxon's Colony Pilot Mine in the Piceance Creek Basin, Colorado has been monitored by periodic pillar stress-determinations End roof-to-floor convergence measurements since 1971. Pillar failure has caused marginal stability in some areas of the mine. Measurements taken during the process of failure provided valuable information on the in-situ pillar strength and long-term behavior of the mine structure. Four pillars were reinforced by 32 mm (1-1/4 in) diameter tensioned grouted bolts to increase long-term stability in the northern area of the mine. Stability of this area is necessary because future plans envision its use as an exit/access and ventilation exhaust. Computer analyses were performed to help evaluate the effect of pillar reinforcement on long-term stability. Results indicate that bolting is particularly useful in stabilizing failed pillars, but its effects are significant only in the immediate area of the pillar. The projected significant decrease in deformation of two extensively failed pillars near a main entry indicates that bolting reinforcement will assist the long-term stability of the opening.

Jan 1, 1985

By Bryan Ulrich

On March 20, 2014, the ?Roundtable Discussion on Innovations in Heap Leaching and Mine Waste Management? was hosted by Knight Piésold and Co. at the Red Lion Hotel in Elko, NV. This topic is timely, and formed something of a lead into the inaugural ?Workshop on Innovations in Waste Rock, Tails and Heap Leach Management.? That workshop, which grew from the 2013 Elko Roundtable, was held on the University of Nevada, Reno (UNR) campus on May 19, 2014. The workshop was hosted by UNR and the University of British Columbia. Knight Piésold?s local representatives, Bryan Ulrich and Logan Jensen, were joined by Tom Kerr, president of Knight Piésold?s, U.S. practice; Rick Frechette, senior vice president, environmental services and Jeff Coffin, geotechnical engineer, from the Denver, CO office for the roundtable event. The event played host to about two dozen attendees from several western states.

By M. Spangler

FirstMiss Gold installed a supervisory control system based on Pavilion Technology's Process Insights neural network software and Oil System's PI data historian. The primary goal of the system was to control the acid preconditioning circuit. However, our first successful control loop was on the limestone feed rate to the neutralization circuit. The information gained by modeling these circuits and the better control provided by the system has resulted in reagent savings of over $80,000 per month.

Jan 1, 1995

By J. C. Judd

The U.S. Department of the Interior, Bureau of Mines investigated solvent extraction methods for recovering gallium from filtrates produced from sulfuric acid leaching of hydrometallurgical zinc residues. These filtrates typically had pH values of 0 to 0.5, too high an acid content for effective gallium recovery by traditional extractants. Of the nine extractants studied in batch tests, 0.3M octyl phenyl acid phosphate (OPAP) in a kerosene diluent was the most effective extractant for gallium recovery, extracting up to 99.9 pct of the gallium. This high extraction was possible when the iron in solution was maintained in the reduced ferrous state. Batch tests also showed that a 1.5M-H2SO4 strip liquor recovered 98 pct of the gallium from the loaded OPAP. A continuous solvent extraction system was designed and operated based on McCabe-Thiele analyses of equilibrium loading and stripping isotherms. When 12 L of leach filtrate containing 0.32 giL Ga and 20 giL Fe were pumped through the system in 32 hr, results showed that 95 pct of the gallium was extracted in four stages at an A:O ratio of 1:1. Stripping the loaded solvent with a 1.5M-H2S04 liquor recovered 94 pct of the gallium in four stages at an A:O ratio of 1:4. A two-stage, 50-pct-phosphoric acid strip successfully removed ferric iron from the OPAP, allowing high gallium extractions to be maintained.

Jan 1, 1990

By C. R. McClung

"Molybdenum, in the form of molybdenite, is a common byproduct of many copper mining and concentrating operations. Contrary to most copper sulfide minerals where recovery is principally liberation controlled, the recovery of molybdenite is much more complex with fully liberated particles commonly lost to tailings. It is this complexity that facilitated the necessity for a geometallurgical investigation of molybdenite from the Bingham Canyon deposit. The aim of this investigation was to determine the effect mineralogy and/or mineralogical attributes (i.e. morphology, angularity, liberation, size) have on the recovery of molybdenite in the processing circuit. A set of samples was collected from the ores and products produced by Rio Tinto Kennecott’s Copperton Concentrator. All samples were analyzed using normal polarized reflective light, Mineral Liberation Analyzer (MLA) and X-ray diffractometery (XRD) techniques. The results of this investigation revealed the presence of two distinctly different types of molybdenite, which have been identified as the two polytypes of molybdenite, i.e. hexagonal (2H) and rhombohedral (3R). The 2H polytype occurs as textbook-shaped particles in quartz-molybdenite veins that are located in the core of the Bingham Canyon deposit. The 3R polytype occurs as disseminated, “ball”-shaped particles with a dull or frosted appearance along the margins of the mineralized intrusive body or central core. Concerning their metallurgical behavior each type exhibits unique metallurgical properties that are consistent with those reported in the published literature. For instance, the 2H polytype is easily ground and kinetically “faster” floating with surface attributes that are amenable to higher rates of recovery. This commonly results in the production of a high quality molybdenum concentrate under normal operating conditions common to most copper concentration operations. In contrast, the 3R polytype is difficult to grind and kinetically “slower” floating with surface attributes that less amenable to recovery under normal operating conditions. Therefore, in deposits with higher concentrations of the 3R polytype, modifications to the normal operating parameters may be necessary to improve the recovery of molybdenum. Finally, a preliminary investigation into the metal budget of porphyry Cu and Mo deposits indicates that a positive correlation may exist between the total economic metal budget and the polytype content. For example, higher concentrations of the 3R polytype appear to be more prevalent in Au-enriched porphyries versus Mo-enriched porphyries. This observation may also serve to help explain why it is easier to obtain high rates of recovery in Mo ± W-rich deposits, which contain little to none of the 3R-polytype. Consequently, the potential for the metal budget of a porphyry deposit to be used as an indication of the molybdenite polytype present, as well as its impact on the economics (exploration) and/or on the processing circuit (plant design) could exist. In conclusion, this investigation highlights the necessity of understanding the mineralogical properties or characteristics of molybdenite and/or other economic mineral(s), as these can and will have a direct impact upon the performance and plant optimization."

Jan 1, 2016

By William R. Nash

Case histories of water problems in underground mine shafts and tunnels are outlined to illustrate varying degrees of sophistication in grouting techniques currently applied in such construction. The interception of water underground can be one of prior knowledge, i. e., the site investigation provides necessary data to plan a grouting program prior to construction start, or excavations may encounter "changed conditions" or unplanned water. The latter includes catastrophic inflows and emergency situations where the structural integrity of a building or concrete shaft lining are jeopardized. The presentation of this paper outlines some applications of grouting work in underground mine construction throughout the United States in a variety of geological provinces and for a variety of minerals.

Jan 1, 1984

By Gösta W. Diehl

During the past five years the use of all-hydraulic drilling equipment has increased rapidly, and we have now reached a point where the hydraulic drill has become an alternative to be seriously considered for every drilling operation. The hydraulic drills have many advantages over pneumatic drills, but they also have certain characteristics, which makes a potential user hesitate before making the switch. And when the decision to use hydraulic drills has been made, the user then faces another choice of equal difficulty: the choice between different hydraulic systems. This paper discusses some of the factors to be taken into account when considering hydraulic drilling equipment for rapid tunneling in hard to medium hard rock. THE CHOICE BETWEEN HYDRAULIC AND PNEUMATIC EQUIPMENT The choice between a hydraulic and a pneumatic drill is much more than just that: It is a choice between two different systems. The choice is usually based on the productivity, economy or ergonomy of the system. However, also other factors should be considered, such as service capabilities and training requirements. And finally the choice will be influenced by the need to acquire know-how in the new technology. Productivity If the prime consideration is to achieve the highest possible tunneling speed, then the high penetration rates of the hydraulic

Jan 1, 1979

By D. Men, R. Q. Honaker, J. C. Hower

"Maceral separation is one possible technique to expand coking coal resources, either by increasing the blending ratio of non- or slightly caking coal or rationally utilizing the coking coal according to the different behaviors of macerals in coking. The success of maceral enrichment for a given coal using flotation is dependent on the differences in flotation kinetic rate among macerals. This paper presents the flotation rates and zeta potentials of the vitrinite-rich and inertiniterich maceral concentrates obtained from six coal samples of different ranks – from high-volatile B bituminous coal with mean random vitrinite reflectance (Rrandom) of 0.64 percent to low-volatile bituminous coal with Rrandom of 1.73 percent – using a float-sink procedure. The flotation results show that vitrinite-rich concentrates can be produced in tap water using froth flotation only with frother for the high-volatile B and A bituminous coal samples, but there was no good performance for the high-volatile C bituminous and medium- to low-volatile bituminous coal samples. The zeta potential results show that froth flotation could probably separate the high-volatile C bituminous coal samples well by adjusting pH, but there was no good performance for the higher-rank coal samples. IntroductionMetallurgical coke is used in blast furnaces for steel production and is produced by the carbonization of coking coal. Because caking coals are rare and expensive, they are often blended with non- or slightly caking coals. Considering the tight supply-demand situation for coking coal in China, it is important and urgent to develop highly efficient technologies for rationally utilizing and expanding coking coal resources (Mochizuki et al., 2013)."

Jan 1, 2015

By S. Frisch

Selenium (Se) in effluent waters has become a significant challenge for a number of industries. Toxicity studies and new reports of occurrence have certainly caused increased scrutiny and actions in the mining industry today (Sandy, DiSante, 2010). While initial activity in addressing selenium-laden waters were conducted in the coal mining industry, today a much broader range of mine types are beginning to search for solutions to selenium treatment of their mining waters, including both hard rock mining operations (gold, nickel, zinc, etc.) and soft rock mining (phosphate, uranium, etc.) Throughout North America, new regulations covering selenium discharge are either in place or being developed. While a number of physical/chemical approaches show promise in selenium treatment, when considering optimal cost and treatment efficacy issues, biological treatment ? particularly fixed-film biological treatment - is reported to be one of the most promising approaches to managing selenium-containing waters. Among biological treatment options, the fluidized bed reactor (FBR) has undergone extensive pilot testing for Se treatment in the last three years, including eight pilot programs, and the first FBR system is now operating at a coal mining site. In all tests, the FBR has been shown to be a highly cost-effective approach to selenium treatment based on its physical and performance characteristics. This is consistent with its history for reaching low levels of treatment in similar service for other oxyanions including perchlorate, nitrate and various metals. With the results from the most recent testing, the technology has been shown to be adaptable to an ever-broadening range of influent water conditions and flow rates, in varying terrains and weather conditions. Selenium occurs in various valence states from -2 to +6 (Sobolewski, 2005). The speciation of selenium plays a critical role in understanding the effectiveness of any approach for removal, especially to low levels. Elemental selenium is insoluble and has little effect on living organisms. Colloidal fractions are known to occur. In aqueous environments such as coal mining runoff, selenium is most often found as the oxygenated anions of selenite (SeO32-, Se4+) and selenate, (SeO42-, Se6+). Selenium is commonly found in mining waters in concentrations ranging from 3 to >12,000 µg/L. The U.S. National Primary Drinking Water Standard MCL is 50 µg/L for selenium (EU 10 µg/L; WHO 40 µg/L). The National Fresh Water Quality Standard is 5 µg/L for selenium. The U.S. Fish and Wildlife Service has recommended that the National Fresh Water Quality Standard be lowered to 2 µg/L to protect fish, waterfowl, and endangered aquatic species. Several states have followed with enforcement actions at these same low levels. In Canada, permits may require stakeholders to monitor levels in water or biota, or to comply with guidelines, at either the National or Provincial levels, which can be as low as 1 or 2 µg/L in surface waters. Treatment of selenium in all mining waters presents a series of challenges. It can be present in relatively dilute concentrations and in streams with variable and often high flow rates. It can be present in many soluble and particulate forms, which may affect process design and treatability. The conditions of influent selenium-containing waters can vary, with issues such as temperature, pH, total dissolved solids and other contaminants affecting the ability to remove selenium selectively and economically. Treatment of selenium often results in the generation of a concentrated by-product requiring disposal, and re-release from residuals can occur. Beyond these issues, the physical environment of a mining operation can offer challenges such as remoteness and harsh and/or cold weather. Currently available selenium treatment technology (Table 1) includes biological treatment and a number of physical-chemical processes ? such as ion exchange, membrane filtration and adsorption ? that vary widely in efficiency and cost. Most treatment options remain either too costly and/or unproven for selenium-containing streams for reasons that include poor removal efficiency, variability in volume and composition of influents, system sizing and logistics considerations, as well as operating requirements.

Feb 23, 2014

By R. J. Hickson

The giant El Abra copper mine project under development in Northern Chile will come on stream in early 1997 at a production rate of 500 million pounds of cathode copper annually. Ownership is 51 % Cyprus Amax and 49% Codelco. Cyprus Amax has construction responsibility. Total project cost is $1050 million. El Abra will incorporate latest operational technology with labor efficient practices to achieve world class productivity levels and low initial cash costs of about US $0.35/lb copper. Development procedures for the almost 800 million ton oxide deposit, as well as the hydrometallurgical processes selected for recovery of the copper values are discussed. Construction costs and pertinent technical data are shown.

Jan 1, 1996

By Robert C. Pease

Siskon Gold has started gold production at the San Juan deposit in northern California. The mine is located about 27 km (17 miles) north of Nevada City, CA (Fig. 1). It is scheduled to be-come the largest underground gold mine in California. It is situated in the western foothills of the Sierra Nevada Mountains in the Mother Lode gold belt. The property consists of 8 km2 (2,000 acres) of patented land. It is situated in a topographically high area known as San Juan Ridge, which is 213 m (700 ft) above and between the Middle Yuba and South Yuba rivers. The elevation is between 790 and 914 m (2,600 and 3,000ft) above sea level. Annual precipitation averages 1,346mm (53 in.). Most of that falls between November and April.

Jan 1, 1996

By Yaxiong Wang, Yong Li, Yuedong Wu, Haigang Dong, Jiachun Zhao

A roasting-leaching process has been developed for concentrating platinum group metals (PGM) from a PGM-containing ferroalloy. The optimized roasting conditions used a NaOH-to-ferroalloy mass ratio of 1.2 at 600 °C for 2 h. The roasted product mainly comprised Na4SiO4, NaFeO2 and Fe2O3. The iron oxide and sodium silicate associated with the roasted product were leached in water and dilute sulfuric acid solution, while the PGM was concentrated in the acid-leached residue. A PGM concentration of 23.4% was obtained with assaying, and the concentration factor of PGM from the ferroalloy to PGM concentration was greater than 9.

Jun 14, 2021

By Jerry Luttman, Mark Dubay, Anvesha Dogra, Elizabeth Dwyre, Wei Hu

The I-75 Modernization Project (Segment 3) just north of Detroit, MI is being delivered as a Design-Build-Finance-Maintain (DBFM), which is a Public Private Partnership (P3). The project includes design and reconstruction of 5.5 miles of freeway, including a depressed section with a 20,630-ft long, 14.5-ft inside diameter, up to 100-ft deep stormwater tunnel, multiple drop shafts, and a pump station. The project also includes a maintenance term of 25-years. This is the first time MDOT has used a P3 delivery method for a major highway and tunnel project. This paper describes the development of the P3 documents, including performance-based tunnel and shaft technical requirements, and the approach to project risk management, financing, and insurance coverage.

Jun 23, 2024

By Yanli Huang, Qiang Sun, Jixiong Zhang, Yuzhuo Jiang, Dan Ma

Mining accidents due to inrushing water are caused by sudden changes in seepage owing to structural damage to key strata of the aquiclude. The backfill mining method can be used to control the movement of the strata, weaken the evolution of fractures owing to damage by key aquiclude strata (KAS), and control the channel for inrushing water. This study develops a mechanical model of a composite beam on a non-uniform elastic foundation of the KAS by analyzing the coupling relationship between the filling body and the key water-resisting layer of the aquiclude. Equations of the deflection-induced deformation, bending moment, and maximum tensile stress of the KAS are analyzed, formulae to calculate the allowable tensile stress and backfill materials’ compaction ratio (BMCR) to ensure structural stability are derived, and are used to provide the critical BMCR for their structural instability. The results of field measurements show that the subsidence of the roof was 600–700 mm with an average BMCR of 80.63%. Damage to the overlying strata was controlled to ensure safe mining and the disposal of solid waste from the aquifer. The research here is useful for ensuring safety during mining while protecting water resources.

Jul 12, 2022

By Ross K. Digby

The Victorian Tunnelling Centre (VTC) has evolved its purpose from purely training workers, to a tunnelling industry education city. The VTC has now focused its activities on the areas of technology; education and training; applied research; and career education. were identied as the hus around which activities would e conducted to create a tunnelling industry education city. The VTC provides ustralias tunnelling industry with an education facility that not only provides laour training, ut enales the industry to come together to meet, participate in product launches, undertake research, work with new technologies and promote tunnelling careers.

Jun 23, 2024

By Robert J. Brison

The intent of this paper is to try to answer the complex question: "How much grinding capacity should be designed into a new mining - milling project?" Much work has been devoted to the technical aspects of grinding, the "how" of grinding. But, in the writers' opinion, the economically important subject of optimizing grinding capacity in mill design has not received the attention it deserves, at least not in published literature. For the purpose of this paper it is necessary to distinguish clearly between optimizing operation and optimizing design. Optimizing operation involves making the best use of the available plant and equipment. Much excellent work has been done and is being done in this area. In operation, one can make tradeoffs of production rate and recovery, or recovery and grade, to maximize profits. However, this must be done within the limitations imposed in the design stage. Flexibility is limited by the existing plant. In project design, on the other hand., there are no preexisting physical limitations except as imposed by nature itself; the planners and designers have considerable latitude in their efforts to optimize the overall project. They have complete control over a large number of important design variables from the mine to the finished product. The grinding process design is one of these variables. It not only determines the capital cost charges for grinding and the maximum grinding capacity that operating management will have to work with, but also determines the range of capability of the auxiliary equipment, which fixes the degree of flexibility possible in operation. This paper is concerned with optimizing design rather than operation. The terms "optimum" and "optimize" are sometimes used in a technical sense in connection with grinding, recovery or grade. In this paper, these terms are used only in the economic sense. In this age of working large deposits of ever lower grade ores, it is in¬creasingly essential to design projects and plants for optimum economics, based on the most applicable economic. criteria. Since grinding represents a large part of both capital and operating cost of the project, and largely determines percentage recovery of metal, it should be included as one of the factors in optimizing the project design. This paper is written primarily with reference to medium to large copper concentrators employing conventional wet grinding. However, the principles involved should be applicable to other ores and other methods of grinding.

Jan 1, 1972