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By W. J. Maloney, H. E. Tuneblom, A. J. Smith

The development of a large regional water supply in Southern West Virginia faced several hurdles. This paper deals with the design and construction of the intake shaft where a combination of cutting edge mechanical excavation technologies were utilized to produce a world class project.

Jan 1, 1999

By J. Boldy

The discovery of the Delbridge massive sulphide deposit in the Noranda district of Quebec resulted from an appreciation of the significance of a mercury leakage halo as a litho-chemical pathfinder to blind ore, coupled to its location relative to a geologically inferred ore-bearing horizon within complex volcanic stratigraphy. Systematic sampling of micro-fractures present in outcrops delineated a diagnostic anomaly, drawing attention to an area considered prospective for buried ore. The massive sulphide deposit was subsequently found to be blind and topped at the 110 metre level.

Jan 1, 1984

Planning the primary development of a mineral deposit usually is done only once during the life of a mine. As a consequence, not optimizing the primary development of a mineral deposit can cause disadvantages over the whole life of a mine and can lead to higher costs. As a contribution to an improved primary development design procedure beyond the already existing work a study of the identification and quantification of the influence of parameters on primary development based on operational data from working mining operations has been completed. Based upon such a quantification of the influence of parameters a sequenced decision model was developed, that enables to develop an initial concept for the design of the accesses from the surface based upon a few critical parameters. Such a model will not replace detailed planning work within a feasibility study, but it is intended to reduce the amount of planning efforts at an early project stage through providing information what are most likely economic solutions, based on simple rules.

Jan 1, 1997

By Juan P. Camus, Sergio G. Jarpa

INTRODUCTION Chuquicamata mine is a large copper open pit located in the north of Chile. It has been in production for more than 80 years and now is in the later stages of its life with declining grades and increasing haul distances. As any old mine with long established organizations and procedures, it has faced one of the most difficult managerial problems: to conduct changes. Since 1994, the new management team has been devoted to a modernization pro- gram focused mainly in restructuring business operations. downsizing the labor force and introducing a mine planning pro- cess consistent with the company's goal. The explicit identification of the latter issue has been made because of the management’s conviction about the importance of it in the profitability of the business. The mine planning process at Chuquicamata mine has evolved from an old manual technique towards a modern computarized system, however the fundamental aspects and the procedures involved in this process have been the same. For in- stance, the cutoff grades policy has not changed during the last said 20 years, regardless of the changes to grade distribution, commodity prices, plant capacities, etc. To review and rethink this process, the strategy was to work with a small group of inside engineers and the management in order to interact with several mining consulting firms. The idea was to open a discussion about the fundamental principles used in the mine planning activities and, at the same time, to explore new ideas to improve the operating practices that have been followed for a long time. Thus, the goal of this paper is to present the mine planning approach that was the result of this work and the improvements that were obtained. THE CHUQUICAMATA MINE The Chuquicamata division is one of the five operations of Codelco and is currently devoted to operating the Chuquicamata mine; a copper open pit, concentrator, smelting and refining complex. It also operates Mina Sur which consists of an open pit mine, a leaching-solvent extraction and electrowinning process. There is also a dump leaching and a Sx-Ew plant for treating the low grade suifide ore (under the cutoff grade) from the Chuquicamata mine. Prior to reviewing the mine planning process, capacities for the Chuquicamata sulfide line were 640 000 tpd of run of mine material, 160 000 tpd of ore and 4 200 tpd of concentrate. This meant a total annual movement of around 230 Mt in the mine and a treatment of 57 Mt in the concentrator with a stripping ratio of approximately 3: 1. The old exploitation strategy consisted of an "open scheme", which means that all benches must have a berm with enough space to allow the traffic of large haulage trucks. This fact implied working with a great flexibility, but the consequences were flat slope angles. One of the reasons for using this operating practice in short term planning was that the grade consistency from the pit has always been a controversial variable. Schedules were designed to meet grade targets accurately and they were modified immediately when deviations become apparent. This reality led to working with a great deal of independent pushbacks and it required flexibility in the traffic system. The cutoff grade policy has been the same for a long time and the figure was 0.5 %CUT either for sulfide or oxide ore. This was the reason why a balance was produced between the mine and the concentrator. The mine is operated principally with benches of 26 m but in the upper part of the pit there are benches of 24 and 30 m. The main production equipment is shovels of 28 and 34 yd3 along with trucks of 200 and 240 st. Regarding the crushing facilities, there are two main systems located inside the pit and an old one in the periphery, close to a stockpile and near the concentrator. In the north part of the pit there is a crusher for waste material and in the north-east area is located the ore crusher facility. Both restrict a significant area of the pit and the mine sequence is considering to move the ore crusher again to the south-east sector where it used to be located. Relating to the waste crusher, the idea is to move it up to a bench near the surface in order to use the same conveyor belt and, at the same time, liberate an important sector. THE MINE PLANNING PROCESS The objective of the mine planning process is to develop a mine plan for the life of the deposit, in harmony and consistent with the company's goal. It includes the origin and destiny of tonnages and grades from the mine(s), along with the necessary resources to carry out this plan such as: fleet requirement, consumable goods, manpower, etc. From the financial point of view, it means to establish the cash flow for each time period of the mine life. That is. the revenue, capital investment and operating cost associated with each period of the plan. The fundamental principles involved in the mine planning pro- cess have been a controversial issue for a long time. One of the reasons for these divergent views could be produced by the difficulty of modeling a complete extraction policy for a geological re- source. This particular problem has to deal with several decision variables that are time dependent. So far, there is no mathematical formulation for managing such a problem, and the approach to solve it is based on a circular analysis combined with certain heuristic tools. So, a global model rather than a formula, requires a clear understanding of mine operations as well as economic theory and its application to a non-renewable resource. In order to manage this complex problem, the mine planning process can be conveniently split off in three stages which are conceptual, feasibility and operational planning. The characteristics of each stage are basically the timeliness in which they are developed and the kind and number of variables involved. Thus, conceptual planning is devoted to the definition of key variables like: the size of the operation, mining method, extraction se-

Jan 1, 1996

By Yuting Xue, Todd Minoski, Daniel Su, Mark Van Dyke, Mark Mazzella, Steve Hicks, Zoheir Khademian, Nicole Evanek

Design of chain pillars in a longwall mine setting is one of the main engineering controls on the global and local stability of the mine. Challenges with roof instabilities, floor heaves, and pillar failures can be alleviated, to some extent, by careful design of gateroad configurations and chain pillars. This work investigates three different alternative designs in a three-entry-system gateroad configuration with an inter-panel barrier in a longwall mine in Virginia. Starting with the current design, which has faced challenges such as excessive floor heave, a geomechanical model developed in 3DEC software is validated by monitoring in-situ pillar stress in the yield, abutment, and barrier pillars in the mine. Next, three different chain pillar designs are evaluated in the model with a focus on roof sagging, floor heave, and pillar average stress. Results show the relationships among the alternative designs and instabilities associated with roof, pillar, and floor in the study mine, providing insights for optimizing longwall mining designs in deep-cover settings.

Feb 1, 2025

By A. E. Sears

The intent of geomorphic landform design is to mimic the function of the natural landscape. In a surface mine reclamation application, geomorphic landform design has the potential to mitigate stream loss. This research quantified features of reference landforms and performed computer modeling of a conceptual valley-fill design using geomorphic landform principles. A design was created using published design parameters with a software design tool (Carlson Natural Regrade with GeoFluv). Three additional designs, incorporating field-obtained regional geomorphic landform parameters, were completed and resulted in a geomorphic landform design specific to southern West Virginia. This work illustrates the importance of field determination of the ridge to head-of-channel input and drainage density parameters. Values of ridge to head-of-channel distance measured in this study were consistently an order of magnitude greater than software recommended values (SRV); SRV = 24 m (80 ft) versus field= 220 m (720 ft). Similarly, for the drainage density parameter, the software recommended range is 6 to 9 km-1 (80-120 ft/ac). The field measurements for central Appalachia quantified the drainage density to range from 5 to 5.3 km-1 (67-70 ft/ac). This affects Appalachian valley-fill stream construction by making the stream lengths shorter and the land slopes steeper with straighter headwater channels compared with areas in the southwestern United States. While the application of geomorphic landform design to surface mine sites presents challenges, this work provides support for future applications.

By Zibo An, Yong Hu, Hao Zhang, Hong Cao, Jun Xue, Zhen Luo

Phosphogypsum (PG) is a by-product of solid waste produced by the phosphate chemical industry. Owing to large emissions and massive storage issues, it causes a potential threat to our environment. Hereby, a method based on the precipitation and crystallization properties of calcium phosphate-magnesium ammonium phosphate (CP-MAP) followed by its adsorption on active solid waste powder is used as the main principle to treat PG harmlessly. The treatment process and its influencing factors for orthogonal experiments along with the associated environmental safety have been studied in detail. The results showed that this method could simultaneously remove soluble phosphorus and soluble fluorine from PG. When ACP, FA, HBM, and Ur were added to PG at a dosage of 3%, 13.5%, 0.04%, and 0.04%, respectively, the residual phosphate and fluoride contents was 0.061 mg/L and 0.856 mg/L respectively in the PG leaching solution. Compared to state-of-the-art approaches, this method exhibited higher efficiency in removing phosphate and fluoride. Most importantly, this method does not cause any secondary pollution in the process, thereby showing a novel application in large-scale industrial treatment.

Aug 29, 2022

By William A. Riggs

Salt, a non-metallic mineral whose use dates back to earliest history of mankind, and whose uses cover the widest range for any mineral raw material, is generally of widespread occurrence in commercially recoverable quantities in the oceans, in certain sedimentary rocks, in certain lakes, and in certain underground waters. Passing over geological detail, it may be said that, generally, all of the salt in the world has, at least once in geological history, been in solution in the oceans. The total salt content of today's oceans, said to equal 4.5 million cubic miles of rock salt, dwarfs the tremendous volume of salt in the sedimentary rocks. In 1958, United States production of salt was about 22 million net tons and world production about 82 million. About 3 percent of United States production is table salt, with chemical, municipal, industrial and agricultural uses taking the remainder. Mississippi River. Figure 1 has been expanded sufficiently to show also those salt occurrences and producing points which might economically ship salt into these states. Although widespread in occurrence, salt deposits and possible production areas generally show very sharp local boundaries. Salt-for non-table use is susceptible to ?place value?, a common problem in low valued raw materials, where transportation costs often exceed the value of the material f.o.b. the production point. Generally only large capacity mechanized salt producing operations can compete in United States markets, as contrasted to the many small local salt evaporators in less industrialized Mexico.

Jan 1, 1962

By M. R. Shahan, C. E. Seaman, S. E. Mischler, T. W. Beck

"Float coal dust is produced by various mining methods, carried by ventilating air and deposited on the floor, roof, and ribs of mine airways. If deposited float dust is re-entrained during a methane explosion, without sufficient inert rock dust quantities, this float coal dust can propagate an explosion throughout mining entries. Consequently, controlling float coal dust is of critical interest to mining operations. Rock dusting—adding inert material to airway surfaces—is the main control technique currently utilized by the coal mining industry to reduce the float coal dust explosion hazard. To assist the industry in reducing this hazard, NIOSH’s Pittsburgh Mining Research Division (PMRD) initiated a project to investigate methods and technologies to reduce float coal dust in underground coal mines through prevention, capture, and suppression prior to deposition. Field characterization studies were performed to determine quantitatively the sources, types, and amounts of dust produced during various coal mining processes. The operations chosen for study were a continuous miner section, a longwall section, and a coal handling facility. For each of these operations, the primary dust sources were confirmed to be the continuous mining machine, longwall shearer, and conveyor belt transfer points, respectively. Respirable and total airborne float dust samples were collected and analyzed for each operation and the ratio of total airborne float coal dust to respirable dust was calculated. During the continuous mining process, the ratio of total airborne float coal dust to respirable dust ranged from 10.3 to 13.8. The ratios measured on the longwall face were between 18.5 and 21.5. The total airborne float coal dust to respirable dust ratio observed during belt transport ranged between 7.5 and 21.8. INTRODUCTION During the extraction of coal, various dusts are produced that can be classified into two categories of interest: respirable and float coal dust. Respirable dust (diameter 10.0 = µm) has the capability to penetrate beyond the terminal bronchioles and become deeply embedded into the human lungs, leading towards respiratory disease complications (Potts et al., 1990). Float coal dust is defined as coal dust having a diameter less than 74 µm (Harris et al., 2009). During the mining process, float coal dust can be deposited on the floor, roof, and ribs of mine airways. With the proper mixture of methane gas (between 5% and 15%) and an ignition source, a methane gas explosion can occur (Barker and Humphreys, 1996). The resulting explosion can develop a pressure wave which lifts the deposited float coal dust material into the air, propagating a disastrous coal mine explosion (Harris et al., 2010). The current technique to safeguard against the propagation of a coal mine explosion is to treat the mine airways with incombustible rock dust (typically limestone), maintaining an 80% total incombustible content of the deposited material. NIOSH initiated a project to investigate control technologies that could be aimed at reducing the amount of float coal dust allowed to deposit in the mine airways using three approaches: prevention, capture, and suppression prior to deposition. Reducing the float coal dust being generated during the mining processes and capturing or suppressing airborne float coal dust prior to deposition would in turn increase the efficiency of the applied rock dust used to prevent mine explosions."

Jan 1, 2017

By S. Cotten

The concept for a Mine Escape Vehicle (MEV) is motivated by the desire to enable personnel to evacuate a mine rapidly as the first and most desirable response to any underground emergency. Under a NIOSH contract, Raytheon UTD developed a conceptual design for a vehicle that coal miners could use to facilitate their rapid exit from an underground coal mine in the case of emergency, including fire or explosion. The vehicle requirements include the capacity to provide life support functions for a number of miners, operate in an oxygen deficient, low or no visibility atmosphere, and travel at speeds faster than miners can walk out of the mine. To develop the MEV concept, three primary tasks were identified and executed. The first task focused on developing a realistic concept of operations (CONOPS) and associated high level vehicle requirements. The second task identified and evaluated currently available equipment and technologies that could be applied to address the CONOPS. The final task developed a design concept that integrated the identified components for the five (5) MEV sub-systems to provide the foundation for future detailed engineering design and prototype construction. The resulting MEV design concept is centered upon retrofitting existing mine personnel carrier chasses with available, mine-worthy equipment capable of addressing the MEV requirements. This approach offers a practical and economic avenue to prototype development and delivery with minimum lead-time and risk. Furthermore, by integrating familiar and available equipment into the design and CONOPS, the resulting Mine Escape Vehicles should also be more readily accepted by the mining industry.

Jan 1, 2010

By C. L. Emery

The geology and early history of the discovery of potash in Saskatchewan as described by Heber Cole (1) and R. V. Thompkins (2) has dealt with the distribution of the salt beds and their potash content. From the many bore holes through the Prairie Evaporites the outlines and contours of the series have been well established. The thickest section of the salt beds lies to the north of the "barren" area and forms a belt running easterly to a point west of Yorkton. The "barren" area is a region where the salt has been redissolved and the strata have slumped down, to fill the voids. This area is of interest from the standpoint of rock mechanics. The potash-rich beds extend both east and west of the region. mentioned and the top of the Prairie Evaporite is at a depth below 3200 feet from the surface near Saskatoon and dips southerly at an average rate of about five feet to ten feet per mile.

Jan 1, 1966

By R. Xiong, A. K. Fallaw, G. Wang, L. R. Moore, S. Dobson, J. Gu

"INTRODUCTION The Central District of the Florida phosphate industry is located in southwest Polk and southeast Hillsborough Counties, with the Southern Extension lying in western Hardee, eastern Manatee and northwest De Soto Counties. (Lawver, 1983) Figure 1 shows the major phosphate formations in Florida. Almost all Florida phosphate production has been in the Central district with additional production from a mine in the Northern District. The Central District is comprised of the Bone Valley Deposit, a relatively thin layer of high grade ore which has been the main source of phosphate mining for decades, and the Hawthorn formation, a thick but low grade formation which covers most of peninsular Florida. (Lawver, 1983) Florida phosphate producers use the Crago process where calcium phosphate is floated with anionic, fatty acid type collectors followed by cationic reverse flotation of silica with amine type collectors. The phosphate matrix is put through various size separation techniques to obtain a typical -28 Mesh (< 700 µm) particle size for fine flotation and a -16 mesh (< 1190 µm) to +28 mesh (> 700 µm) particle size for coarse flotation. In some cases this coarse fraction can be sent to shipping as an intermediate product. The +16 mesh (>1190 µm) to – ¾“ (19 mm) fraction is sent to shipping as a final product called pebble. Figure 2 demonstrates a simplified diagram of the mining and processing of calcium phosphate (matrix) through a Central Florida mine. Particle size and mineralogy are important considerations in the Crago process for two main reasons, the mass of the particle and the particle surface area. Both of which are dependent on the coordination of the anionic collector to available Ca2+ sites on the minerals’ surface. The increased surface area, and available Ca2+, of the smaller apatite particles will allow for more fatty acid molecules to coordinate than the larger particles, with a lower availability of Ca2+. So, the larger particle would require more collector to be introduced in the system to sufficiently cover the apatite particles to produce a surface sufficiently hydrophobic to attach to air bubbles allowing flotation to occur. (Guan, 2009) The Bone Valley deposit has very favorable flotation characteristics when compared to other mining ores around the world. The matrix of the Bone Valley is relatively high grade, BPL (Bone Phosphate in Lime) values typically range from 6-18%, BPL recoveries, from such a matrix, can often be achieved above 85+% with fatty acid based collectors dosed below 1 #/ton. As mentioned, the sizing is an important step as there is also high grade apatite in the +16 mesh (1mm) that is screened to a final product grade, 60-65% BPL. The lack of quartz in the larger size fraction has allowed the Central District to produce high quality, low cost phosphate product, called pebble, through effective screening with no further beneficiation. (Guan, 2009)"

Jan 1, 2015

By T. N. Andersen

The history of gold and silver analyses shows an interesting duality. On one hand, classical chemical and instrumental methods have evolved to meet changing needs, but on the other hand, the age-old and somewhat empirical fire assay has remained a popular tool. Inasmuch as gold and silver generally have been present in ores at low concentra¬tions, methods have been challenged by a need for sensitivity and selectivity coupled with reasonable assay time. A progressive increase in the importance of low grade materials as ore sources has led to the need for improved methods. Assays of these materials are required for process control and accounting in operating plants as well as for guiding research and development activities. The analytical response to these needs is evidenced by numerous reviews published through 1977-¬e.g., cf Kolthoff and Elving (1966), Beamish and Van Loon (1966, 1972, and 1977), Young (1971) , Mallett (1970), and Sen Gupta (1973). In this paper we present a brief sketch of the history and status of various gold and silver methods. A more detailed summary of atomic absorption spectroscopy (AAS) is Given, as this method has gained in popularity more than other assay methods. AAS methods currently in use are summarized and discussed briefly. This paper is not meant to be a detailed review, as are the above mentioned publications. Rather, we borrow from those works and more recent papers to report trends in gold and silver analyses (particularly in AAS). Also, we discuss some selected poten¬tial problems, and problems that have been experienced by this laboratory.

Jan 1, 1981

By Samuel Frimpong

The preparation of graduate talent for the mining industry is a challenging process, given the multifaceted and complex array of decisions mining engineers make in discharging their responsibilities. As leaders, mining engineers must manage operations, lead project teams, mentor junior engineers, chart the technological directions of companies and deal with several organizations for maintaining the technical and economic feasibility of operations within strategic and tactical framework. Traditionally, the completion of the bachelor of science degree curriculum, with a minimum cumulative grade point average (CGPA), has been used as a graduation measure by Missouri University of Science and Technology (Missouri S&T). Companies have long realized the limitations of the CGPA for recruiting graduates because mining engineering is an applied engineering discipline. Company recruiters have sought to correct the limitations by factoring experience and behaviors into the selection process when recruiting knowledgeable and well-rounded graduates. In this paper, the authors introduce a new metric, which combines the current CGPA (knowledge) with experience and attribute (KEA) metrics to gauge the job readiness of graduating engineers. The KEA metric, also called the job-readiness factor (JRF), is the sum of K (50%), E (30%) and A (20%). The KEA process has been applied to a graduating class in Missouri S&T to demonstrate the value of the new metric system. The results show a fairly positive correlation between CGPA and the KEA metric, with a coefficient of determination of about 70%. Overall, the KEA metric system provides a better picture of the capacity and job readiness of graduates entering the workforce.

Jun 1, 2013

By E. D. Attanasi, M. W. Bultman

Most statistical measures of mineral exploration continued downward in 1986. Prospects in 1987, however, appear to have brightened for exploration as the overall economic health of the mineral industry improved. The Society of Economic Geologists (SEG) survey of US and Canadian firms documents the marked decline in US mineral exploration since the early 1980&apos;s. From 1981 through 1986, the total exploration expenditure by US firms declined by more than 73% (Fig. 1). Exploration expenditures by US firms declined by 66%%. US based professional exploration staffs were slashed by 70% (Fig. 2). Also, many firms have written off some reserves because of weak markets or have not delineated new reserves to replace production. The Canadian government reported that between 1981 and 1986 Canadian molybdenum re¬serves had declined by 30%, lead and zinc reserves by 20%, and nickel and copper reserves by 15%. The SEG survey also shows the immediate impact of the decline in the dollar on US exploration expenditures abroad. Between 1981 and 1986 the share of the exploration dollar spent by US firms outside of the country declined by nearly one-half and exploration acreage abroad declined by 69%. The downward trend of the last six years may finally be near the end as a leaner and more efficient minerals industry starts to gear up for a period of increased activity. An industry survey reported to Mine Development Bimonthly (Metals Economics Group) indicates that mining firms were expected to increase exploration expenditures in 1987 by at least 10% above 1986 levels. Worldwide expenditures for 1987 by companies covered by the report were expected to top $800 million. It is the first time since 1984 that an increase in exploration expenditures was expected. Also, the US Bureau of Land Management reported a 30% increase in new mining claims for fiscal year 1987 over fiscal year 1986. The bulk of the increase in claims was in states having intense gold exploration, particularly Washington and Oregon. Much of the optimism is undoubtedly due to the increases in metal prices that occurred in 1987. Some of these increases include gold, average price up about 20%; copper, average price up 24%; aluminum, average price up 26%; and nickel, average spot price up 20%. The SEG survey shows precious and base metals accounting for 89% of the exploration expenditures in the US in 1986, up 52% from 1981. The US Bureau of Mines estimates that 80% of US exploration activity in 1986 was directed entirely at gold (1986 Minerals Yearbook). Worldwide exploration plans indicate the continued dominance of gold exploration. New programs were announced for Ghana, Malaysia, and Indonesia. Major new exploration programs were also launched for lead-zinc sulfides in Morocco, platinum and palladium in northwest Brazil, rare earth minerals in Greenland, diamonds in Indonesia, and heavy minerals off-shore Australia. Significant gold discoveries were made at Angels Camp in Oregon; South Africa, near Potchestroom; India, in the Chitradgra district; Northern Ireland near Mt. Speerin; Chile, San Cris-

Jan 5, 1988

By P. James

"INTRODUCTION Dissolved salts build up in leach solutions over time as leach solutions mature. Iron can be particularly problematic for hydrometallurgical copper production as it can compete with copper for separation during solvent extraction (SX), leading to iron bleed through and reduced electrowinning (EW) efficiency. In its ferric form (Fe+3) it precipitates as a hydroxide at fairly low pH and can promote clogging and disrupted flow within leach pads as a result of localized pH deviations – such as might result from weather events. The solids formed can cause relatively irreversible degradation to pad productivity. The precipitation pH point for solids formation falls as the ferric (Fe+3) concentration increases, making the issue more prevalent as leach circuits for ore bodies containing appreciable iron get older. It can also be more of a concern when leaching lower grade ore like tailings where lower copper content may increase iron/copper ratios which may increase iron accumulation rates. The precipitation pH point for solids formation falls as the ferric (Fe+3) concentration increases, making the issue more prevalent as leach circuits for ore bodies containing appreciable iron get older. It can also be more of a concern when leaching lower grade ore like tailings where lower copper content may increase iron/copper ratios which may increase iron accumulation rates. Control of ferric (Fe+3) levels could be used to lower ferric (Fe+3) levels and keep the ferric hydroxide [Fe(OH)3] precipitation pH sufficiently high to avoid the associated solids formation and potential leach pad occlusion. Control of ferric (Fe+3) and ferrous (Fe+2) tenors and ratios may also provide benefit for improving sulfide ore leaching or bioleaching performance and effectiveness. Traditionally options for such control were rather limited with bleeding contaminants off or chemical dosing treatments being prominent methods. Now Blue Planet Strategies introduces a new option. Blue Planet Strategies (BPS) focuses on improved resource utilization targeting electrolytic treatment and waste conversion into resources and has now demonstrated that its proprietary DeMet™ (Dynamic Electrolytic Metal Effluent Treatment) platform is highly effective at efficient low-cost conversion and control of iron content in leach streams in addition to targeted dilute metal recovery.[1-6] This powerful new capability provides a new treatment opportunity to control leach solution ferric (Fe+3) and ferrous (Fe+2) tenors to enhance leach operation production."

Jan 1, 2015

By M. L. Smith

SAG mill throughput is often a bottleneck in increasing metal productivity. Improved size distribution control in the mill feed can be used to increase mill production. This can be accomplished by controlling blast fragmentation as a function of energy input and rock characteristics. Such an approach requires the integration of rock fragmentation and comminution modeling. The optimal mill feed size distribution is determined by adapting a population balance comminution model to a measure of rock hardness or fragility which is also relevant to the rock fragmentation that is achieved by blasting. The relation between the size distribution and fragility of the feed and of the mill&apos;s product and the critical size in a SAG mill is examined.

Jan 1, 1993

By William F. Lawrence

Within ten to fifteen years, coal-fired power generating stations will be producing over 40 million tons of coal-ash annually in the United States. The largest part of this ash will be emitted from the boilers, entrained in the off-gas stream and trapped as flyash. However, about 15 percent (six million tons) will he removed from furnace bottoms as bottom ash or molten coal-ash slag. In current practice, the molten coal-ash slap is generally water quenched and then transported either as a slurry or in granular form to storage or disposal areas. Although a small amount of this slag is marketed as abrasive grit or as roofing granules, most of it is considered a waste material which must be disposed of at considerable cost to the power companies and therefore to the public. The accom¬panying economic loss in terms of unused raw materials and displaced land is not as obvious but is just as real. The Coal Research Bureau at West Virginia University under the sponsorship of the U. S. Bureau of Mines has been examining the possible utilization of both coal-ash slag and flyash in the mineral wool industry as a method of partially alleviating the growing national problem of solid waste disposal. The potential marketing situation for mineral wool manufactured from coal-ash is illustrated in Figure 1. This bar graph depicts the growth in value of mineral wool shipments1 from 1949 through 1966. In this 17-year period, the value in¬creased by over 350 percent, more than 20 percent per year, despite a decrease in unit cost of mineral wool. It therefore appears likely that coal-ash could find a steady market as a supplement to existing sources of raw materials for mineral wool production.

Jan 1, 1970

Water contact angles in the air/water/francolite, fuel oil/water/francolite and fuel oil/water/quartz systems were measured to describe the wetting characteristics of these minerals. It was found that white francolite is more hydrophilic than black and tan francolite both in the absence and presence of fatty acid addition. The contact angle measurement results do not correlate very well with the flotation response and it appears that kinetic aspects of the bubble attachment process play an important role in the francolite-fatty acid flotation system.

Jan 1, 1997

By C. M. Brown

The semi-autogenous grinding circuit at the Utah Mines Limited, Island Copper Operation, located on Vancouver Island, British Columbia, is described. Since start-up of the operation in 1971 grinding circuit throughput has been progressively increased from the initial design rate of 30,000 to the current average 45,000 tonnes per day (t/d) of porphyry copper ore. Key developments and modifications in operating and maintenance practice that have contributed to this increase are reviewed.

Jan 1, 1985