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By Guowen Dong, Qiyong Li, Qiang Li, Tiejun Chen, Zhihong Wu, Junying Wan

In order to search for a more economical and effective roasting additive for extracting vanadium from stone coal and solve the question of the disposal of PVC waste, PVC waste was proposed as a novel roasting additive to enhance extraction of vanadium from stone coal and realize the comprehensive utilization of resources. By adding this roasting additive, the leaching rate of vanadium reached 94.5% under the conditions that the PVC waste addition was 10 wt%, the roasting temperature was 800 °C, the roasting time was 1 h, the sulfuric acid concentration was 15 vol%, the liquid-to-solid ratio was 1.5 cm3 /g, the leaching temperature was 95 °C, and leaching time was 4 h, which increased by 8.4% compared with that of blank roasting. The thermodynamics calculation showed that the decomposition reaction of mica was performed easier with the participation of HCl produced by pyrolysis of PVC. Meanwhile, The SEM-EDS and aperture test analyses demonstrated that more cracks and voids appeared in the roasted sample by roasting with PVC waste, which could facilitate the release and extraction of vanadium. Additionally, the TG-DSC analyses indicated that PVC could provide large amounts of heat for the roasting process to promote the decarbonization and roasting efficiency of stone coal because of its high calorific value.

By Guowen Dong, Qiyong Li, Qiang Li, Tiejun Chen, Zhihong Wu, Junying Wan

In order to search for a more economical and effective roasting additive for extracting vanadium from stone coal and solve the question of the disposal of PVC waste, PVC waste was proposed as a novel roasting additive to enhance extraction of vanadium from stone coal and realize the comprehensive utilization of resources. By adding this roasting additive, the leaching rate of vanadium reached 94.5% under the conditions that the PVC waste addition was 10 wt%, the roasting temperature was 800 °C, the roasting time was 1 h, the sulfuric acid concentration was 15 vol%, the liquid-to-solid ratio was 1.5 cm3 /g, the leaching temperature was 95 °C, and leaching time was 4 h, which increased by 8.4% compared with that of blank roasting. The thermodynamics calculation showed that the decomposition reaction of mica was performed easier with the participation of HCl produced by pyrolysis of PVC. Meanwhile, The SEM-EDS and aperture test analyses demonstrated that more cracks and voids appeared in the roasted sample by roasting with PVC waste, which could facilitate the release and extraction of vanadium. Additionally, the TG-DSC analyses indicated that PVC could provide large amounts of heat for the roasting process to promote the decarbonization and roasting efficiency of stone coal because of its high calorific value.

Paper 1: Risk of gastrointestinal cancers from inhalation and ingestion of asbestos, John F. Gamble Selikoff suggested the hypothesis that asbestos exposure is causally associated with increased risk of gastrointestinal (GI) cancers in an early study of insulation workers. In this paper, John Gamble reports on his extensive review and reanalysis studies that address the relationship between asbestos exposure and GI cancers. The purpose of the review was to evaluate the association between asbestos exposure and risks of stomach, colorectal, colon, and rectal cancers separately. Gamble evaluated exposure?response (E?R) relationships using surrogate exposure methods ??because few individual studies of GI cancer have reported exposure? response analyses.?? The principal asbestos exposure surrogates are: (1) the rate ratio (RR) for lung cancer; and (2) the percent of mesothelioma among exposed workers. Analyzing SMRs for GI cancers from individual studies versus the risk of lung cancer and mesothelioma provides crude estimates of E?Rs for asbestos exposure and GI cancer. Briefly, Gamble justifies the surrogate approach as follows:

Jan 1, 2007

By Steve Kral

The mining industry has waited a long time for the resurgence in commodity prices and demand for its products that it is now experiencing. Base and precious metals prices have reached levels not seen before. Coal prices have also shot up. And there are encouraging signs in the uranium market. China’s nearly out-of-control economic growth is said to be the leading driver for much of this resurgence. But demand for energy, metals and building materials in the United States and elsewhere are also major factors. The result is that expansions and new mines for nearly every commodity are in the works.

Jan 1, 2005

By Manuel Pérez-Tello, María M. Salazar-Campoy, Óscar Rodríguez-Hoyos

A common problem during the operation of the Kaldo furnace is the formation of accretions and blockage of the injection lance due to splashing. This typically occurs during the processing of anodic slimes for the production of doré. In this work, a 1:7 scale model of the Kaldo furnace currently in operation at Metalúrgica de Cobre, Mexico, was built in transparent acrylic plastic. Water and colored oil were used to represent the doré and slag, respectively. The model was used to determine the maximum penetration distance of the injection lance under a variety of experimental conditions at room temperature. Input variables included two types of oil, rotation speed, inclination angle, and the fraction of the vessel volume occupied by the liquids (φ). A mathematical correlation for the dimensionless maximum penetration distance of the injection lance d ∗ as a function of φ, the Froude number of the liquids in the vessel (Fr), the Reynolds number of the gas jet (Re), and the oil-to-water kinematic viscosity ratio (η)was developed. Reasonable agreement between the calculated and experimental values was obtained. Digital images showed that splashing was caused by the turbulent gas jet impinging upon the free surface of the liquids, whereas the cascade effect was negligible. The results of this work have been useful to improve the operation of an industrial Kaldo reactor.

By Manuel Pérez-Tello, María M. Salazar-Campoy, Óscar Rodríguez-Hoyos

A common problem during the operation of the Kaldo furnace is the formation of accretions and blockage of the injection lance due to splashing. This typically occurs during the processing of anodic slimes for the production of doré. In this work, a 1:7 scale model of the Kaldo furnace currently in operation at Metalúrgica de Cobre, Mexico, was built in transparent acrylic plastic. Water and colored oil were used to represent the doré and slag, respectively. The model was used to determine the maximum penetration distance of the injection lance under a variety of experimental conditions at room temperature. Input variables included two types of oil, rotation speed, inclination angle, and the fraction of the vessel volume occupied by the liquids (φ). A mathematical correlation for the dimensionless maximum penetration distance of the injection lance d ∗ as a function of φ, the Froude number of the liquids in the vessel (Fr), the Reynolds number of the gas jet (Re), and the oil-to-water kinematic viscosity ratio (η)was developed. Reasonable agreement between the calculated and experimental values was obtained. Digital images showed that splashing was caused by the turbulent gas jet impinging upon the free surface of the liquids, whereas the cascade effect was negligible. The results of this work have been useful to improve the operation of an industrial Kaldo reactor.

By E. C. Robertson

D.G. Wilder I found the suggestion that the amount of displacement of a fault can be numerically related to the thickness of gouge or breccia to be both intuitively satisfying and intriguing. I have long agreed that there is some type of relationship between the amount of gouge and the amount of displacement of faults. I congratulate the author for developing a numerical relationship between them. However, I am concerned that the limits for applying this relationship be fully understood. An underlying assumption in this approach is that there is either a uniform thickness of gouge or breccia along a given fault or the thickness does not vary widely. Since it is not always possible to confirm this, the displacements derived by this method should be viewed with caution unless significant fault extent can be observed. At the Nevada Test Site, in drifts constructed in granite for test emplacement of spent nuclear reactor fuel, we found a fault with 0.3 to 0.4 m (12 to 16 in.) of clay gouge. Within a few meters of this location, the fault had no clay gouge, but rather consisted of a highly fractured zone with significantly altered rock and some slickensides. Based on Fig. 1, the 0.3 to 0.4 m (12 to 16 in.) thickness of gouge would indicate a displacement in excess of 30 m (98 ft). However, no gouge thickness would indicate essentially no displacement. Based on a quartz vein that terminated on the fault, and is not identified nearby, an estimated displacement of more than a few meters was made. This estimate is consistent with that obtained using the regression line proposed in the paper if the 0.3 to 0.4 m (12 to 16 in.) thickness for the gouge is used. However, using the regression curves with zero thickness would not yield results consistent with what was observed in the field. Therefore, it is important to recognize that the suggested procedure would properly yield a range of probable displacements. ? *Work performed under the auspices of the US Department of Energy by the Lawrence Livermore National Laboratory under Contract W-7405-Eng-48. Reply by E.C. Robertson It is certainly true that the t (thickness) of gg-bx (gouge and breccia) on a fault does vary along the fault. My observations have been that near the termination of a fault, the displacement d is small and the t is also small, whereas the maximum d and t will usually be found in the central part of the fault. The information on gg-bx and t of the fault found in granite in the NTS tunnel by Mr. Wilder could be interpreted somewhat differently than he does. He speaks of the fault changing within a few meters from 0.3 to 0.4 m (12 to 16 in.) of clay gg to "a highly fractured zone with significantly altered rock and some slickensides," but no gg. The highly fractured rock may be taken to be bx, rock not so finely ground as gg but still crushed by the fault movement, equivalent to the gg in my usage, and probably occupying about the same t. Mr. Wilder's estimates for the fault in the NTS tunnel for t of 0.3 to 0.4 m (12 to 16 in.) and for d of a quartz vein, in excess of "a few meters," would place the point on the low side of the central trend line in my Fig. 1, at the lower limit. There is, of course, a problem with determining d using displacement of only one planar surface. It would be greater or lesser depending on the rake of the movement. Finally, estimating the d of a fault from its t should be made with awareness of our present uncertainties, as pointed out by Mr. Wilder. Although the central trend line in my Fig. 1 has a ratio of d/t of 100, I have put the limiting ratios at 10 and 1000. Understanding of the values of the ratio will be improved only with collection of more data, for which the discussion of Mr. Wilder is much appreciated. G.C. Waterman E.C. Robertson's paper provides significant information to a geologist attempting to deduce fault offset by noting the products of structural dislocation. However, considerable mapping in underground and open-pit mines, and examination of structures produced in different geological settings, have convinced me that gouge and breccia thickness are controlled by geological conditions and fault movement. The following paragraphs suggest geological variables that control them. 1. Depth of Loading A near-surface fault resulting from tensional stress has more breccia/gouge than is produced by a similar stress at considerable depth. A deep-loaded compressional stress may produce a linear zone of schist, or structural dislocation may occur along an earlier formed belt of schist. Such "shear zones" are common in Canadian mines in precambrian rocks. In neither case can offset be directly deduced by an analysis of the minimal gouge/breccia in the shistose rocks. At greater depth, stress may be partially to wholly relieved by flowage. I vividly recall first noting the regional "Midas Thrust" in the Lark mine, Bingham Mining District, UT (where we called the structure the North Fault). My recorded notes, as I remember them, showed a narrow gouge streak separated the `Jordan" and "Commercial" limestone units from impure, muddy limestone beds of uncertain stratigraphic position. The visible structure did not indicate the great importance of this premineral fault

Jan 12, 1984

By Richard P. Stahura

During the next five years, there will be a trend toward performing belt cleaner maintenance while the conveyor is in operation. Most mines and industrial sites that use belt conveyors have an in-house safety rule that the belt must be stopped before maintenance work can be performed on any belt-cleaning device. This is a good rule. Most cleaning devices are positioned against the belt with hardware that requires the service technician to reach or climb into the chute to change the cleaning blades. But this simple rule becomes more complicated as production demands require higher conveyor availability and provide fewer conveyor outages. In many cases, today’s conveyors are expected to be available 98% to100% of the time.

Jan 1, 1996

By Zhigao Xu, Ming Wu, Jun Qu, Rong Zhang, Wenrui Nie, Ruan Chi, Zhenyue Zhang, Zhengyan He

The leaching processes and mechanisms of weathered crust elution-deposited rare earth (RE) ores with ammonium sulfate were investigated for their high efficiency and low consumption exploitation. The leaching process is a typical reversible heterogeneous ion exchange reaction in liquid–solid systems. The raw ore and leached tailings were analyzed by SEM and FTIR. Results indicated that no dissolution or structural changes of the ores occurred during the leaching process. The migrations of RE3+, Al3+, and NHþ 4 in the leaching process were studied under various liquid–solid ratios. With increasing liquid–solid ratio, the content of NHþ 4 adsorbed on clay mineral by physical effect decreased, NHþ 4 adsorbed by ion-exchange increased, and the content of exchanged RE and Al increased until reaching equilibrium. The adsorption capacity of the RE ores for NHþ 4 remained about 0.61–0.63 mg/g. It was difficult to desorb NHþ 4 from RE ores with washing water alone. The leaching mechanism of Al was the same as RE, but the leaching of Al slightly lagged behind RE. At the optimal liquid–solid ratio of 0.4:1, the leaching efficiencies of RE and Al were 95.51% and 90.99%, respectively, and 56.65% of NHþ 4 remained in the leachate. After impurity removal and RE precipitation, the leachate can be reused by replenishing fresh leaching agent. The aim of this study is to enhance the leaching process of RE, restrain the leaching of impurities, and decrease the consumption of leaching agents.

By Zhigao Xu, Ming Wu, Jun Qu, Rong Zhang, Wenrui Nie, Ruan Chi, Zhenyue Zhang, Zhengyan He

The leaching processes and mechanisms of weathered crust elution-deposited rare earth (RE) ores with ammonium sulfate were investigated for their high efficiency and low consumption exploitation. The leaching process is a typical reversible heterogeneous ion exchange reaction in liquid–solid systems. The raw ore and leached tailings were analyzed by SEM and FTIR. Results indicated that no dissolution or structural changes of the ores occurred during the leaching process. The migrations of RE3+, Al3+, and NHþ 4 in the leaching process were studied under various liquid–solid ratios. With increasing liquid–solid ratio, the content of NHþ 4 adsorbed on clay mineral by physical effect decreased, NHþ 4 adsorbed by ion-exchange increased, and the content of exchanged RE and Al increased until reaching equilibrium. The adsorption capacity of the RE ores for NHþ 4 remained about 0.61–0.63 mg/g. It was difficult to desorb NHþ 4 from RE ores with washing water alone. The leaching mechanism of Al was the same as RE, but the leaching of Al slightly lagged behind RE. At the optimal liquid–solid ratio of 0.4:1, the leaching efficiencies of RE and Al were 95.51% and 90.99%, respectively, and 56.65% of NHþ 4 remained in the leachate. After impurity removal and RE precipitation, the leachate can be reused by replenishing fresh leaching agent. The aim of this study is to enhance the leaching process of RE, restrain the leaching of impurities, and decrease the consumption of leaching agents.

The design/optimization of the pump and cyclones to maximize grinding circuit efficiency first requires a concise understanding of, and clearly defined specifications for, desired grinding circuit performance. Subsequently, the pump and cyclone performances that provide the desired grinding circuit performance can be defined. Once the performance requirements of each process unit are determined, selection of physical equipment parameters that will deliver those performances can be carried out. The Functional Performance Equation shows that there are two distinct efficiencies in closed circuit ball milling. One of these, ?Circuit Classification System Efficiency? (CSE), is the fraction of ?coarse? material inside the ball mill upon which grinding energy is deliberately expended, versus the remaining fraction of ?fines? or finished size material inside the mill, upon which grinding energy is wasted (on over-grinding). The role of the pump and cyclones is to maximize CSE, which is the desired grinding circuit performance. The obtainable, target CSE for any given circuit can be estimated, for example by referring to a plant data base of CSE?s. Design/optimization guidelines then allow us to specify the pump and cyclone performances that will achieve the target CSE. This paper outlines the step-by-step, systematic method of pump and cyclone selection which maximizes CSE. Numerous examples of its successful application are provided.

Jan 1, 2011

By J. Wickens

Haile Gold Mine is located near Kershaw, South Carolina, approximately 50 miles south of Charlotte, North Carolina. Gold was first discovered at the site in 1827, and the property has been through several operating periods since. The relatively low grade and refractory nature of the deposit has been a challenge for efficient processing since the start of mining on the property. Early processing techniques included gravity separation and barrel chlorination. The last operation employed heap leaching in 1985 through 1992. After being closed and reclaimed for over 20 years, the property is poised to re-open and produce once again. In this chapter of the Mine?s legacy, flotation and ultra-fine grinding will be employed to achieve economic extraction of gold and silver from the pyrite. This paper briefly reviews some of the mine?s history, the metallurgical testing, the flowsheet and discusses design considerations. KEY WORDS: process plant, flash flotation, leaching, SMD, stirred media detritor

Feb 27, 2013

By C. Geier

Our national interests are best served by having well-trained first responders available for rapid response to emergency situations in underground mines. However, recent mine disasters have indicated that focused training on hazard recognition, decision making, leadership, and incident command center protocols needs to be incorporated into a comprehensive training plan for all personnel involved in a mine rescue. Currently, there are very few training opportunities for the mining professionals responsible for making the crucial decisions needed during a mine emergency. With recent technical advances, it is now feasible to use a computer-based simulator to train mine rescue personnel for mine emergencies in their mine or any other mine at a convenient location without interrupting normal mining operations. This simulated environment teaches underground search and rescue procedures, communications, and decision making based on real-time information and hazard recognition. Since 2010, the Colorado School of Mines (CSM) has offered computer simulated mine rescue training using an incident command center (ICC) directed towards enhancing decision making and communication skills. To support this training, CSM partnered with Rite Solutions, Inc. to modify, enhance, and create simulation software for mine rescue applications. The focus of this partnership is to modify existing simulation technology to address the unique nature of underground rescue operations, make it applicable to any underground operation, create generic scenarios, and develop customized computer simulations, including scenarios, for specific sites. Need The CSM mine rescue computer simulator helps mine rescue teams train for mine disasters where they would be deployed for the rescue and recovery of a mine. It provides additional, readily accessible opportunities to train without affecting mining operations. While the simulator lacks the physical demands of a rescue situation, it is very effective for teaching mine rescue procedures as rescuers seamlessly walk through the process and work out issues before going underground. The mine rescue simulator also introduces students to coordination through an ICC which stresses the use of effective communication and hones decision making skills. Most of the mine rescue teams participating in this study had never trained with an ICC controlling the actions of the mine rescue team making this an extremely valuable training opportunity. In addition, the mine rescue computer simulator allows instructors to modify mine rescue scenarios and tailor the scenarios to each team?s individual needs resulting in a more versatile training experience. The goal is to ensure that all teams receive the maximum benefit from each training session.

Feb 27, 2013

By Taekhee Lee, Steven Mischler, Shawn Vanderslice

Diesel-powered equipment is widely used in the underground mining industry and their popularity results from properties including efficiency, versatility, reliability, and durability. In the U.S. approximately 7,700 diesel engines were used in 177 underground metal and nonmetal mines as of 2005 [MSHA 2005], and these numbers have likely increased since that time. Diesel engines have been shown to be a major contributor to submicron carbonaceous, respirable, and total particulate mass in the air of underground metal mines [Zielinska et al. 2002, McDonald et al. 2003], and their extensive use results in approximately 13,000 underground metal/nonmetal (M/NM) miners (MSHA 2005) in the U.S. being potentially exposed to aerosols and gases emitted by diesel engines (MSHA 2001, MSHA 2006). Exposure to diesel exhaust has been linked to adverse health outcomes including cancer, cardiovascular, and respiratory diseases (Attfield et al. 2012, Silverman et al. 2012), and in 2012 diesel exhaust was categorized by the International Agency for Research on Cancer (IARC) as a Group 1 human carcinogen (IARC 2012). Exposure to diesel particulate matter (DPM) is especially concerning for underground miners since underground mine environments have been shown to have some of the highest levels of diesel exhaust in the U.S. (EPA 2002, MSHA 2001, MSHA 2006, Pronk et al. 2009). Due to the potential for elevated DPM concentrations in mines, the Mine Safety and Health Administration (MSHA) promulgated a rule to limit exposures of metal/ nonmetal underground miners to DPM to an eight-hour time-weighted average (TWA) of 160 μg/m3 total carbon (TC) (MSHA 2001, MSHA 2006, MSHA 2008). Since this rule went into effect, DPM exposures have been reduced, but are still elevated when compared to other occupations (Pronk et al. 2009, Noll et al. 2015). MSHA compliance data between 2005 to the present show that underground blasters represent 21% of all DPM overexposures in M/NM mines and are one of the highest exposed professions in mining often resulting from low ventilation in the area where blasters are working. Because mines have difficulty controlling DPM in these low ventilation areas, additional control technologies may be needed to reduce exposures. Administrative controls, where miners work on off schedules or upstream of diesel vehicles to avoid the exposures to DPM (Noll et al. 2015) is one possibility, but these types of solutions are not always feasible or practical. A canopy air curtain is another potential control technology to help reduce exposures of blasters to DPM. Listak and Beck (2012) showed that this control technology reduced respirable dust concentration under a roof bolter’s canopy by 67%–75% and recommended air velocities greater than 0.5 m/s for dust reductions of greater than 50%. Additional work showed that a canopy air curtain could be designed for a shuttle car, and some initial testing by the National Institute for Occupational Safety and Health (NIOSH) showed reductions of respirable mine dust between 66% to 70%. As seen in Figure 1, the canopy air curtain delivers clean air over the operator’s breathing zone. A fan draws in air through a filter to capture the dust and then supplies clean air beneath the canopy where a miner is working. The use of the canopy air curtain to reduce exposures to diesel particulate matter was first discussed by Noll et al. (2020). In this study the diesel canopy air curtain (DCAC) reduced the DPM concentrations under the canopy by up to 90%. For this research the DCAC was designed to attach to the basket cover of an ammonium nitrate fuel oil ( ANFO) loader (Figure 2) and provide clean air blowing over the miners as they work in the basket. The initial tests used clean air drawn from the mine’s intake and did not try to filter the DPM from the air. Filtering DPM has extra challenges when compared to filtering mine dust particles. The particles of DPM are smaller (submicron and nanometer) than dust particles (greater than 1 micron); therefore, the filtration system must be adjusted to capture submicron particles. A MERV 13 filter was used in previous canopy air curtain research on mine dust control, but this filter is only designed to capture 50%–75% of submicron particles, and this capture efficiency is too low for removing DPM, so a higher efficiency filter is needed. Higher efficiency filters increase the pressure across the filter media which results in decreased airflow or leaks around the filter and reduces the effectiveness of the control for protecting miners from DPM. The optimal filter needs to capture DPM particles at a high efficiency while still allowing the needed airflows to prevent contaminated air from entering the miner’s breathing zone. Listak and Beck (2012) recommended velocities greater than 0.5 m/s for dust reductions of greater than 50%. However, airflows too high have been shown to reduce miners’ thermal comfort; Roghanchi et al. (2016) suggest the optimal velocity for thermal comfort is between 1–2 m/s. Thus, airflow over the miner should be limited to the range of 0.5 to 2.0 m/s. This current study presents the results from research evaluating the reduction of DPM concentrations, under the DCAC, when using a higher efficiency filter (MERV 16) to remove DPM from the DCAC airstream.

Feb 1, 2024

By W. G. Pariseau

"BACKGROUND The Crandall Canyon Mine (CCM) is located in the Wasatch coal field in central Utah. Extensive longwall mining at CCM preceded a decision to recover coal left in barrier pillars on each side of a set of main entries extending far into the mine. During mining in March, 2007, a major bump occurred that led to cessation of mining in the barrier pillar on the north side of the mains and to begin mining in the south barrier pillar. Subsequent loss of ground control in August, 2007, through coal pillar bursts had tragic consequences with loss of six miners and shortly afterwards three rescue team members. An MSHA report describes the events in detail (MSHA, 2008). An important lesson to be learned from CCM is the need to select a computer program suitable for pillar design. Computer programs are engineering tools; selecting the right tool for the job is essential. In this regard, a minimum program requirement is capability for computing stress distribution within a pillar and adjacent roof and floor strata. The reason is simple: strength of rock is stress dependent; both vary from point to point in pillars and adjacent strata. Preferably, one should follow the evolution of stress as the mining plan is executed from start to finish to assure safety or to warn of a potentially hazardous situation developing as mining progresses. The approach to the problem posed by barrier pillar mining at CCM is conventional in the sense of applying well known finite element modeling techniques to analysis of a proposed mining plan. The engineering purpose is to evaluate the mining plan with respect to ground control questions including roof, floor and pillar safety. Results of previous two dimensional finite element analysis indicated pillar safety was by no means assured (1, 2, 3). Geometry of barrier pillar mining is limited in two-dimensions, for example, crosscuts are not seen in vertical cross sections. While three dimensional analyses are preferred, even moderately sized regions often lead to very large problem sizes that exceed computer capabilities. However, a recent development in finite element modeling of tabular deposits allows for whole mine analysis in three-dimensions. This technique involves a dual node – dual mesh concept. Both have been used separately in much earlier work, but are only now are used in combination. Estimation of subsidence at a kilometer scale and details of stress, strain and displacement about barrier pillars, main entries and pillars, and so on at a meter scale are possible (4). No compromise involving empirical formulas or similar guess work is required. Indeed, all calculations are based on first principles: physical laws, kinematics, and material laws. Effects of joints and in situ variability or uncertainty in strata properties, elastic moduli and strengths, may be included for greater realism. An important objective of the present study is to assess the role of three-dimensional pillar geometry in arriving at conclusions regarding barrier pillar mining versus two-dimensional results."

Jan 1, 2015

By O. Schumacher

Guidelines for publication of information defining the amount and grade of resource material available in a deposit are well established in Australasia?s JORC Guidelines, Canada?s National Instrument 43-101 and similar documents in South Africa, Peru, Chile and elsewhere (Table 1). These guidelines provide standards to apply regarding the geologic and geostatistical basis for describing the level of confidence to be assumed for an estimate of the amount and grade of material. All of them provide for a strict distinction between material that has been demonstrated to be economically mineable and that which has not, and all of them provide terms such as ?resource? or ?exploration information? to be applied to data regarding material that has not been demonstrated to be economically mineable. It is not until an economic study has been completed that we are allowed to apply the powerful terms ?reserves? and ?ore? to our deposit material. Indeed, the application of these eye-catching labels immediately tells the reader that an economic study has been completed by a qualified person, and that the results of the study were positive. While extensive guidelines are established for describing the level of confidence in published geologic data, and the nature of the economic study that should be accomplished before using the terms ?reserves? or ?ore?, surprisingly little guidance is provided regarding the process to use, and the level of detail to attain, for estimating capital and operating costs for the study. The intent of this paper is to lay the groundwork for development of suitable guidelines for the cost estimating process. While the codes do not provide specific instructions for the cost estimate, they do provide numerous hints and clues to help determine the level of detail expected and the type of procedure to apply in preparing the estimate. This paper will first review the various codes to gain an insight into the thinking of the drafters of the codes regarding their unstated expectations for the cost estimating process. Based on this review, a series of guidelines are proposed for this purpose.

Feb 27, 2013

By Russell Barr

Strategic alignment throughout a mining organization is critical to effective strategic execution and good business performance. The alignment of strategic goals and objectives in the different operations and functions in a large global operation is a challenge facing executive management in light of legacy organizational structure and processes. An additional confounding factor in strategic alignment of acquisition and new project expansion is that mining projects and operations are frequently multi-cultural in terms of country origin as well as organizational culture. This brings together many cultures to accomplish the project or complete an acquisition (Financial Times, 2006). The mining industry presents numerous challenges to executives and mine management by the nature of the operation and organizational processes used in running it. For example, there seems to be an inherent conflict between the individual goals of the components that constitute a mineral processing operation (the mine, the mill, the smelter and the refinery), and lowest cost production of the final product. Managers of these component operations have specific objectives passed down from the corporate executives. Those site managers tend to be compensated based on throughput and perform to this main metric. This tendency to maximize throughput for an individual operation can contradict the general process flow of a best-fit solution to producing the final product at the lowest cost. In an effort to become more aligned, firms are increasingly using balanced scorecard derivatives at great cost and effort (whether internally or externally sourced) in an attempt to help with strategic alignment. Balanced score-cards have been effective in providing a good strategic scorecard to companies with some measure of strategic alignment. Many firms make this effort of using balanced scorecards. However, most firms are not strategically aligned or, at best, are sub-optimized, in terms of strategic alignment (Kaplan and Norton, 2006). The tendency in mining operations is to focus on operating individual components as optimally as possible and stretching existing assets to the best possible production rates. ?Making tons? is the common operational focus. Numerous factors underpin this focus and can affect the ability of the operation?s ability to align properly with the corporate strategy and strategic goals. Critical factors are legacy organizational structure, culture, processes and performance measurement systems. Changing anyone of these factors to affect better strategic alignment requires significant attention to managing change.

Jan 1, 2008

By J. L. Cupp, R. C. Ford, B. R. Romig

The Climax Molybdenum Mine, located near Leadville, Colorado, is the site of a lengthy mining history spanning more than 80 years. In the 1960’s, extraction of molybdenum from oxide ore located adjacent to the massive molybdenite sulfide deposit resulted in the construction of an earthen core dam to impound fine-grained oxide tailing in the Eagle River Valley. Through recognized value of water storage and reclamation opportunities, a tailing removal project was initiated in 1993 to convert the impoundment facilities to a post-mining beneficial land use of developed water resources. An evaluation of the effect residual materials and lake dynamics would have on in-stream water quality was performed. Eagle Park Reservoir stands as a model for future reclamation efforts that involve water delivery to highly sensitive receiving waters. This paper provides a case study on project development, the evolution of water quality assessment, and the regulatory framework that contributed to this project’s success.

Jan 1, 1999

By J. S. Meyer

Although salmonid fishes (including Pacific salmon and trout) have only intermediate sensitivity to copper (Cu) for traditional toxicity endpoints (survival, growth, and reproduction) and thus do not ?drive? the freshwater Cu criteria (e.g., see species-sensitivity ranking in US EPA 2007), Cu concentrations as low as 1 to 2 µg/L have been reported to impair olfaction (the sense of smell) in salmonids in laboratory experiments using dilute fresh water (see references in Meyer and Adams 2010 and DeForest et al. 2011). Olfaction is crucial to predator avoidance, recognition of kin, and reproductive synchronization (Baldwin et al. 2003); and Cu is of general ecotoxicological concern because it can be present in acid/mine rock drainage, industrial and municipal effluents, agricultural runoff, and urban stormwater runoff (Meyer et al. 2007). Therefore, there is public concern that the current regulation of Cu in surface waters is not stringent enough to protect salmonid fishes against olfactory impairment (e.g., OSU 2007; Pearson 2007). Because the toxicity of metals to aquatic organisms is a function of water chemistry [e.g., acute toxicity of Cu generally decreases as pH, water hardness (Ca2+ + Mg2+), alkalinity, and dissolved organic carbon (DOC) concentrations increase; Meyer et al. 2007], hardness-based freshwater criteria for metals are lower at lower water hardness. The recently-developed biotic ligand model (BLM) explicitly takes into account all four of those water chemistry parameters and the concentrations of other major inorganic cations (Na+, K+) and anions (Cl-, SO42-) to predict the acute toxicity of cationic metals (e.g., Ag, Cd, Cu, Ni, Pb, Zn) to aquatic organisms; and the BLM has been adopted by the U.S. Environmental Protection Agency (US EPA) to calculate acute and chronic national freshwater criteria for Cu (US EPA 2007). Consistent with observations of Cu toxicity to aquatic organisms, BLM-based Cu criteria generally are lower at lower pH, water hardness, alkalinity, and DOC concentrations. Therefore, although threshold concentrations for olfactory impairment by Cu can be low in dilute fresh waters, the corresponding hardness-based and BLM-based Cu criteria for the same waters also tend to be low. This leads to a logical question: Are Cu criteria already protective against impairment of olfaction in salmonid fishes, thus obviating the need to decrease those criteria even lower to protect against adverse effects to this physiological/ecotoxicological endpoint? Herein, we demonstrate that in all studies conducted to date, the US EPA?s BLM-based freshwater Cu criteria would have been protective against olfactory impairment in salmonid fishes, but the US EPA?s hardness-based criteria would not always have been protective. This distinction is important because all 50 States in the USA still use hardness-based Cu criteria in their water quality standards, and only some of the States allow use of the BLM to calculate site-specific Cu criteria. Additionally, in the absence of empirical results for calibration of a saltwater Cu-olfactory model, we demonstrate that a proposed unified freshwater-saltwater Cu-olfactory BLM predicts the US EPA?s current saltwater Cu criteria would be protective against olfactory impairment caused by Cu in sea water.

Feb 27, 2013

By A. L. Williamson, G. Savci

Hydrologic assessments of existing and proposed waste rock stockpiles (WRS) are typically required for application of operating and closure permits at mine sites. The results of these assessments can be used to determine the applicability of the state groundwater permit program requirements to the operation and closure of these stockpiles. The objective of this hydrologic assessment was to ascertain whether or not the WRS at the Ajo Mine have a reasonable probability of discharging potential pollutants to the underlying aquifer under the operational mine conditions. The hydrologic assessment was performed based on an evaluation of climatic data, prediction of the net infiltration rate through the WRS and prediction of additional hydrologic factors such as the time required to deplete available moisture storage in WRS. The predicted net infiltration rates ranging from 0.13 inches/year (1.1x10-8 cm/s) to 0.24 inches/year (1.9x10-8 cm/s) are below the threshold value of less than 10-7 cm/s recommended by Hutchison and Ellison (1992) as being characteristic of conditions that produce no or negligible quantities of seepage. In addition, the total time required to deplete available moisture storage within the Ajo Mine WRS and for any potential seepage from the WRS to reach the underlying aquifer is estimated to range from 260 years to 1,950 years. Collectively, the climatological and geotechnical investigations and the hydrologic predictions demonstrate that the WRS at the Ajo Mine hydrologically do not have a reasonable probability of discharging pollutants to the underlying aquifer. As a result of this prediction, the stockpile will not be regulated under the state groundwater permit program and operational and reclamation requirements will be limited to managing surface water runoff.

Jan 1, 2002