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By Syd S. Peng

Is the laboratory sized coal specimen much stronger than underground coal pillars? The answer to this question has a profound implication to many coal operators. It will decide how much of the coal reserve can be recovered. Or, in many cases, it will determine whether it can be mined economically. Recent trends in several eastern states to establish regulations regarding coal pillar design to prevent surface subsidence long after mining heighten the debates. And, in some cases, these trends highlight the existing practices that seriously question the conventional wisdom. Currently, nearly all researchers and practitioners agree that as the size of rock specimens increases, its strength decreases. This concept has been universally adopted as the basic principle for coal pillar design in the United States. Several formulas have been derived to account for this effect. The reduction factor for underground coal pillars ranges from one-half to one-sixth of the strength obtained in the laboratory, depending on the formula selected. Recent case studies indicate that the reduction factor is not justifiable, at east in some cases. The use of a reduction factor is based on the assumption hat smaller rock specimens as tested in he laboratory contain none or fewer fractures or planes of weakness than larger specimens or underground coal pillars. In other words, the larger the rock materials, the more fractures it contains and subsequently weaker in strength. Do all underground coal pillars have more fractures than laboratory sized coal specimens? Not necessarily. All coal contains cleats. Some are con¬spicuous but others are not. In terms of cleat density (number of cleats per unit area), there is no difference between an underground coal pillar and a laboratory sized specimen, provided the specimen size is larger than the cleat spacing, which is mostly 25 to 75 mm (1 to 3 in.).

Jan 1, 1993

By S. Agarwal

The mining industry collects a significant amount of operational data. However, gleaning useful information from the terabytes of data is difficult, and not just because of the sheer volume of the data. Therefore, an automated tool was developed at the University of Alaska Fairbanks to go through data and apply very sophisticated statistical and neural network techniques, to identify the data streams that are important to a process. This paper presents results from the tool as applied to SAG mill data from a gold mine. The results were benched against commercial tool. Results indicate that there was little or no loss in performance by automating the very complicated process of neural network modeling. Therefore, the intent of the exercise, to examine if complicated modeling tasks can be automated, was realized.

Jan 1, 2011

By V. R. Degner

The trend in both the metallic and non-metallic mining industry is toward processing increasing y higher tonnages of lower grade ores, thereby resulting in a growing interest in progressively larger volume individual machines in the flotation circuit. Because of the relatively high cost of large flotation machine development and evaluation, it is important to establish a rational basis by which the existing smaller flotation machine experience can be extrapolated to establish the configuration of the next larger family of flotation machines with a minimum of development modification in the field. This paper reviews the R&D approach used in translating smaller flotation machine experience to establish the configuration of a 1000 cu ft (2B.3 cu m) induced air machine. Internal cell hydrodynamic principles, which governed this extrapolation, and cell (liquid) surface and froth management considerations will be reviewed, and prototype machine mill tests will be summarized.

Jan 1, 1978

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 William G. Doepken

With most of the mining industry emerging from the deepest business recession in 50 years, primary attention must be given to surviving in a business climate where dominant foreign competition, depressed prices, and potential overcapacity hang heavy over much of the industry. While already shackled with regulatory constraints, restrictive work rules, and high labor costs, mine operators are faced with the difficult task of reducing costs and improving productivity while operating at less than full capacity. To counter these nearly over-whelming factors, the mining industry must continue at the national level to seek regulatory relief and a rational policy toward foreign competition from government-controlled mines. In addition, dialogue with labor organizations must continue to minimize the impact of labor costs, cost escalators, and antiquated work practices. Although national and international actions have a tremendous impact on the future of domestic mining operations, local actions by mine operators to dramatically improve mining productivity will strengthen the industry's competitive stance. Before any substantial progress can be made, however, mine operators must recognize and define areas that depress productivity and establish goals and measures for improvement. Once problem areas are defined, a combination of technological and motivational concepts can be applied to improve productivity. Timing of Improvements While there is an immediate need for productivity improvements, only certain concepts offer near-term improvement potential while others must, of necessity, fit into a mid-range or long-range time frame. Near-term solutions (immediate to one year) likely will involve motivational concepts - work incentives and training - coupled with available technological concepts that require little capitalized development or site testing. Mid-range solutions (one to three years), by contrast, will include technological improvements in mining equipment, techniques, and systems automation. These concepts require a longer period of capitalized development and site testing before use. Long-range solutions (three years or more), which include adoption of new mining methods and computerized control systems, will appear only after prolonged research and capitalized development programs. Motivational Improvements Developing motivational con-

Jan 11, 1983

By W. Randall, Robert M. Leary

Twin 1250 m (4100 ft.) long highway tunnels through Cumberland Mountain near Middlesboro, Kentucky are planned to be completed in the early 1990's. The tunnels will be excavated in sandstones, limestones, and shales, dipping at 35 to 50 degrees and striking nearly perpendicular to the tunnel alignment. In 1985-86 a pilot tunnel was excavated as a crown drift along the alignment of the southbound tunnel. The benefits of the pilot tunnel in reducing uncertainties for main tunnel construction are discussed. The cost-benefits of the pilot tunnel are examined in view of estimated main tunnel costs.

Jan 1, 1987

By H. Haselhuhn

Filtration of hematite ore slurries can be a bottleneck in some hematite processing plants due to inadequate filtration rates caused by the dispersion of fine particles. Flocculation of fine particles in the slurry decreases the specific cake resistance of the filter cake allowing for increased filtration rates. CO2 sparging of hematite slurry was shown to increase the filtration rate by altering the pH and the surface chemistry of the particles. It was shown that an increase in filtration rate of 278% can be accomplished by lowering the pH from 11 to 6.5 using CO2. This was attributed to increased flocculation of hematite and silica particles in the ore.

Jan 1, 2012

By O. MacIntosh

The Round Mountain gold mine, located 90 km (55 miles) north of Tonopah, NV began large-scale production in 1977. During 2008 Round Mountain Gold Corp. (RMGC) relocated the primary crusher facility approximately 300 m (1,000 ft) to make room for the next phase of openpit mining activities. This facility was installed in 1988 and used a log-crib retaining wall and dump pocket. RMGC selected Hilfiker Retaining Walls and Collins Engineers, Inc. to provide design and materials for a new 17-m- (56-ft-) high mechanically-stabilized embankment (MSE) wall to support the haul trucks at the new location. The construction contractor was EarthTech West Ltd. of Norwood, CO. An MSE wall with a cast-in-place concrete face was chosen as a durable, yet less expensive alternative to the reinforced concrete dump pocket that was originally designed for the project. The arrangement of the MSE wall is of an unusual configuration to meet a variety of operational and maintenance requirements. This paper outlines the selection process, design and construction of the MSE wall.

Jan 1, 2010

By V. T. McLemore

In-situ leaching (ISL) operations have been proposed to recover uranium from sandstone-uranium deposits in NM. ISL (also known as in-situ recovery, solution mining, solution-leach mining, leach mining) consists of injecting a leaching solution into the mineralized zone through injection wells. The leach solution migrates through the formation, dissolves uranium and is recovered through production wells. The uranium is then precipitated through a conventional uranium recovery system. Several criteria must be met in order for in-situ leaching to be successful, most important is that the ore body must be amenable to chemical leaching. A number of ISL operations have been conducted in NM in the past (Mobil, Crownpoint; UNC-Teton, Section 23; Grace Nuclear, Hook?s Ranch, Seboyeta, Church Rock; Anaconda, Windwhip). Kerr-McGee (later Quivera Mining Co., Rio Algum) successfully produced uranium from mine-water recovery (recirculated mine water) from the Ambrosia Lake mines (mid-1960s to 2002). Despite some difficulties, ISL is a viable alternative to recovering uranium and could be utilized in the Grants district with proper aquifer characterization and monitoring.

Jan 1, 2009

By J. Herhold, S. Plaum

The focus within the economical optimization of projects, which are conducted underground, shifts more and more towards the conduction of construction work underground. Underground operations can be defined in a wide range of various projects, from underground mining tasks to the construction of engineering structures. Within these projects the conduction of construction work cannot only be defined as absolutely necessary, but sometimes also as optimizing for underground processes (e. g. installation of workshops underground to support the mining processes). Even though the conduction of underground construction might cause positive outcomes concerning the productivity, it also tends to be challenging due to local (sub-surface) conditions and influences. To minimize the risk of failure and sudden events within in the conduction of construction work underground, the tool “work preparation” shall be used. Under the given constraints, such as local influences, expected construction work and set time frame, the conduction of the construction work will be planned in advance by consideration of the construction logistics as well as the selection of a fitted construction technique. PROBLEM DEFINITION Economic Challenges - Raw Material Extraction Enterprises in the construction, mining and raw material extraction industries are increasingly under the pressure to perform economically to sustain their position on the global market. Profitability, efficiency and productivity are the basic conditions to be an active participant on the global market as well as a business partner and contractor for mining and construction projects or projects related to the extraction of raw materials. The completion of large-scale as well as common construction projects above ground has been under the pressure to minimize costs and time whilst maximizing the potential outcome, in the meanings of quality and economic gain, for more than a decade now. Whereas the construction industry has to fight the competition on the market with efficient project outcomes, the mining industry has to find ways to perform efficiently to further conduct the extraction of decreasing raw material deposits. The ongoing exploitation of existing raw material deposits to satisfy the constantly increasing raw material demands world-wide, led to the predicament that economically mineable deposits have decreased dramatically. Even tough mineral deposits do still exist on a global level, near-surface deposits decrease rapidly. Therefore, the extraction of raw materials within mining operations will be relocated further away from the surface in greater depths. With the growing distance from the surface to the mineral deposit, the extraction costs will increase many times over. With that in mind, the extraction costs will increase whilst the economic efficiency will decrease. The extraction of certain deposits will no longer be profitable with the common approach. To further extract raw materials efficiently even in greater depth, companies have to change their practice in different ways. That leads to the fact, if the deposits are in the depths, where the extraction processes won´t be economically, other (support) processes have to be in order to guaranty an economical corporate process itself. This precept will divert the focus from the actual mining process to support processes, where potential for optimization still can be found, as shown in figure 1.

Jan 1, 2019

By M. Gobla

"Since 1994 the Bureau of Reclamation has used risk analysis to evaluate dams. Risk considers both the likelihood of failure and the expected consequences should failure occur. Risk analysis relies on the detailed formulation and critical analysis of all possible potential failure modes. Potential failure modes are formulated by describing the sequence of events that must occur to result in failure, where failure is defined as the uncontrolled release of the impounded reservoir. Failure occurs in response to loadings such as water seepage through a dam (static forces), shaking from an earthquake (dynamic forces), sudden flood inflows (hydrologic forces), or accidental overfilling (operational errors). Reclamation has applied risk analysis methods to inactive and abandoned mines to evaluate tailings dams and flooded underground mines. Case histories of potential failure mode analysis as applied to mine impounded water illustrates what critical technical information is needed to determine if failure is likely, what the consequences of failure are likely to be, and what are the highest risks that should be addressed first. INTRODUCTION Impounded water, and high water content wastes, at mining and industrial processing sites are potential threats to public safety and the environment. Recent failures such as the TVA Kingston flyash impoundment failure in Tennessee in 2008 (5.4 million yd3 released), the Mt Poly Tailings Dam failure in British Columbia in 2014 (5.8 million yd3 released), and the Smarco Fundao Tailings Dam failure in Brazil in 2015 (78 million yd3 released) has renewed attention to this issue. Potential risks from impounded water at mining and industrial sites originates not only from tailing dams, but also from water in process ponds and conveyance pipelines, sediment retention dams, water retention dams, mine pit lakes, and flooded underground mine workings. Conventional water retaining dams also have a record of failure. Notable examples of failure include Vajont Dam in Italy in 1963 (2,600 lives lost), Banqiao and Shimantan Dams in China in 1975 (171,000 lives lost), and Machhu-II Dam in India (10,000 lives lost). After the 1976 Teton Dam failure in Idaho, the Bureau of Reclamation (BOR) established a Dam Safety Program. By 1984 the agency developed and began using risk analysis methods. Reclamation operates and maintains an inventory of hundreds of aging dams which were designed and constructed over a period of one hundred years. Many of these facilities were constructed using engineering design and construction methods which would not be considered good practice according to present day standards. Diligent observational monitoring, detailed site investigations, and the application of risk evaluation to the BOR dam inventory has identified many deficiencies resulting in significant demands on available funding for remediation efforts. The application of risk analysis has allowed the agency to rank sites by risk. By giving priority of remedial efforts to high risk sites, significant improvements to dam safety have been achieved. BOR has also applied risk analysis to evaluate and reclaim inactive and abandoned mining and industrial sites. The application of risk analysis methods in the mining industry is not new, the methodology was transferred to the mining industry more than 15 years ago as discussed in ICOLD Bulleting 121 (ICOLD, 2001), yet failures persist. This paper is meant to convey the lessons learned from BOR’s use of risk analyses in its work on mine sites."

Jan 1, 2017

Current design guidelines for the surface catalysed oxidation of Fe(II) on Fe(OH)3 (Ochre) in passive mine water treatment systems are based largely on empirical data as the details of rates and mechanisms have not fully been revealed. Also the effects of elevated carbonate concentrations, typical of many UK minewaters, on the surface catalysed oxidation rate are unclear. In order to move away from empirical design and towards sound engineering that is based on geochemical and hydrological understanding it is necessary to quantify these reaction rates. In this study the effects of Fe(OH)3 and carbonate on the rate of Fe(II) oxidation by dissolved oxygen have been investigated with an aim to delineate the reaction mechanism. Experiments were conducted under well-controlled conditions. Preliminary results indicate that the surface catalysed reaction is an order of magnitude faster at pH ~6.4 with 200 mg/l Fe(OH)3 than the homogeneous rate and is dominated by an elevated rate of removal and oxidation of Fe(II) from the solution phase. It is also apparent that the oxidation rate of actual adsorbed Fe(II) is substantially retarded with respect to the homogeneous rate. Carbonates were not shown to affect the surface rate within the bounds of this preliminary study.

Jan 1, 2006

The management of risks associated with slope instability is an essential process in the safe and economic operation of open cut mines. The ?slope stability radar? (SSR) has been developed to better manage those risks. The SSR remotely scans rock slopes to continuously measure any surface movement and can be used to detect and alert users of wall movements with sub-millimeter precision. The high level of movement precision and broad area coverage of the rock face can allow for a better understanding of the geomechanics of slope deformation, including magnitude of potential failures and additional warning time of impending instability. Additionally, radar waves adequately penetrate through rain, dust and smoke to give reliable measurements, 24 hours a day. SSR systems have been deployed in many metalliferous and coal mines throughout the world. Over 80 systems are currently being used with current operations in 17 countries. Greater than 200 rock slope collapses and waste dump failures (from several to millions of tones) have been monitored, with ?warning? precursor movements recorded by the SSR. This technology enables a radical change in the management of risks in open cut mining. This paper will present recent case studies of the SSR operation in coal and metalliferous mines. Slope Stability Hazards and Rock Slope Monitoring in Open Pit Mines Surface mining often creates excavation boundaries which are steeper than those would be created under natural processes in the same rock type. They are also undertake quite rapid stress changes during mining and are influenced by significant external forces such as blast vibrations and changes in groundwater. These conditions are likely to allow some degree of instability to develop in the pit slopes. The hazards associated with this instability of pit slopes can be extremely variable in character; ranging from small cobble sized rocks dislodging and falling from mine benches to complete collapse of entire pit walls. It is usually the position of mine geotechnical or geological staff to assess these hazards for mine operations. The assessment of hazards due to unstable slopes can be done by visual observation of precursor signs such as fracturing or rilling of loose material. A more reliable indicator of instability involves the quantitative measurement of outward movement and acceleration of material as an instability mechanism develops. There is strong evidence that small precursor movements of a rock wall occur for an extended period prior to rock slope collapse (Hoek and Bray, 1981). Development of a monitoring system, adopting acceptable slope deformation criteria coupled with warning systems and design of stabilization or risk reduction measures if appropriate has become a standard method of dealing with slope instability. Framework for Slope Hazard Management using the SSR A framework for the management of slope hazards is shown in Figure 1 (discussed in detail in Harries, 2007). Although this framework is specifically for use with the Slope Stability Radar, a similar framework could be developed for other monitoring technologies for monitoring slope deformation. An important point to consider is the sequential nature of the hazard management framework from selecting the context ? identifying hazards ? analysing hazards ? evaluating hazards ? treating hazards. [ ] Once the context of the system has been developed (for example ? to monitor for rapid bench scale collapse), the identification of hazards is the next stage in hazard management. Identification of slope hazards can be relatively straightforward due to the visual nature of the information that the SSR system produces. The areas of excessive wall deformation are highlighted in color maps of the mine slope with a co-aligned photographic image to review the actual site of interest. The responsible personnel must then analyses this data, usually reviewing the deformation style and rate, as well as the size of area that appears to be failing. This information can then be evaluated against organisation protocols (for example it may be only a small sub-bench failure that will be ?caught? by existing benches and hence can be categorized as a lower risk). Once these hazards are evaluated they can then be treated if required. Two major options are available to reduce the risk during this treatment option. We can reduce the likelihood of failure ? for example by buttressing the failure or de-pressurising of water pressures in a slope. However, the most likely plan of action is to reduce the consequence of the failure. One of the most successful opportunities to reduce the consequence of a slope failure is to ensure that no personnel or mining equipment is present at the location of a slope failure when collapse finally occurs. This has been the major benefit of the SSR system to the different mining operations around the world. SLOPE STABILITY RADAR TECHNOLOGY OUTLINE Slope Stability Radar system specifications and capabilities: ? Measurement precision +/- 0.1mm (std.) ? Perforated dish and other features enable it to operate in high wind conditions ? Measurement range is 1700 metres ? Scans 270 degrees horizontally and 120 degrees vertically ? 1 ? 30 minute repeat rate for area scans (varies according to area being scanned ? RF exposure (<0.05mW/cm2) is a factor of 200 below the acceptable limit for the general population (1 mW/cm2) ? Tolerant to vibration and mining equipment ? The system weighs approximately 1500 Kg ? Remote area power supply enabling unmanned, continuous 24/7 monitoring ? Real time monitoring

Jan 1, 2009

By J. C. Wynn

J.R. Hillebrand From the abstract, "A realistic ore deposition model to serve as a conceptual framework." Further, on page 246, "The first area of study was the Josephine peridotite, now identified as an ophiolite terrane, located in northern California and south-western Oregon (Fig. 3). Here, it was felt that we could test the models shown in Figs. 1 and 2 and, if successful, expand the model and test it in other chromite districts in the US and around the world." My comments refer only to chromite deposits in Europe and the Middle East. First, I question that may chromite deposits occur in the idealized ophiolite section shown in Fig. 1. Second, although some podiform deposits (referring to morphology) have been important (Guleman, Turkey, for example), the stratiform deposits are of greater importance in tonnage and frequently grade. Examples include Nada, Yugoslavia; Bulqiza, Albania; Xerolivado, Greece; Saranovsk, Russia; and Abdasht, Iran. Reply by J.C. Wynn Mr. Hillebrand raises two issues that deserve a serious response. One issue is basically subjective and the other is ultimately political. First, the question of where (and how many) podiform chromite deposits are found in the idealized ophiolite section of Dickey (1975). The Dickey model is still perhaps somewhat controversial (Panayoutou, 1980), but our experience in the Josephine ophiolite of northern California (Wynn, 1980; and Albers, 1983) shows that the model works well there. Virtually all the chromite is found in dunite lenses in the tectonite peridotite (harzburgite). Since there are few complete ophiolites exposed in the world, the model will probably never satisfy everyone. The degree of acceptance is a function of any individual&apos;s subjective experience. Second, it is well known that stratiform chromite contributes more than 95% of the chromite reserves presently known in the world, and it could be reasonably argued that there is no point in exploring for or exploiting podiform deposits further. The title of the paper, however, was in part &apos;strategic minerals geophysical research&apos;; chromite is considered a strategic and critical mineral of the US and many other industrialized nations. Stratiform deposits in South Africa and Zimbabwe account for most of the chromite reserves in the world. Unfortunately, experience in modern times with substantial price fluctuations and even resource cartels dictates the wisdom of identifying and maintaining a national chromite reserve, irrespective of the cost. An improved exploration methodology is an essential element in this strategic objective. References Albers, J.P., 1983, "Geological and Geophysical Studies of Chromite Deposits in the Josephine peridotite, California and Oregon," US Geological Survey Bulletin 1546 A-D, 118 pp. Dickey, J.S. Jr., 1975, "A Hypothesis of Origin for Podiform Chromite Deposits," Geochimica et Cosmochimica Acta, Vol. 39, pp. 1061.1074. Panayiotou, 1980, Proceedings, International Ophiolite Symposium, Cyprus, 1979, 781 pp. Wynn, J.C., 1980, "Exploration Geophysical Methods and Strategies, and Their use to Locate and Assess Small Mineral Deposits," Chapter 12 of Meyer and Carmen, eds., The Future of Small Scale Mining, McGraw-Hill, New York, pp. 75-83.

Jan 12, 1983

By J. R. Kunkel, V. Lishnevsky

At mine sites, the operational and required storage volumes of solution and water storage ponds and reservoirs are usually calculated based upon a water-balance approach. Many climatological variables including precipitation, air temperature, solar radiation, pan evaporation and relative humidity along with operating parameters, are used in the water-balance model. The water-balance model often seeks to find a “critical period” of wet or dry climatological sequences which result in a “maximum” pond storage volume or water demand. The problem with this approach is that the designer usually does not know when the project will come on line in relation to the historical record. To avoid this pitfall, we use a more robust approach. In the more robust method, the water-balance calculations begin in each year of the historic climatological record. This method results in the same number of water-balance runs (or realizations) as the number of years of climatological record. Results of this type of analysis guarantee that the critical period or maximum required pond volume will be found if it is assumed that the historical climatological record repeats itself or if a synthetic time series is used. In addition, this method permits calculation of the likelihood (probability) that a given pond volume will be equaled or exceeded in any given month of the operational life of the facility. In this paper, the results of the sizing of a heap leach pad solution pond are presented along with the probabilities of a given solution pond size being equaled or exceeded based on the robust water-balance approach.

Jan 1, 2002

By B. Boberg

Annual requirements of uranium to fuel the 441 currently operating nuclear reactors in the world is 170 million pounds U3O8 (65,385 tonnes U). 2005 global production was 108 million pounds (41,595 t U), of which only 2.7 million pounds U3O8 (1,039 t U), was produced from US uranium deposits. US nuclear utilities represent 26% of the global nuclear capacity but only obtain 2.5% of their supply from US uranium mines. Since 2002 annual global production increased from 93 million pounds U3O8, a 15% increase and the US production increased from 2.3 million pounds to 2.7 million pounds, a 17% increase. Reasonably Assured Resources (RAR) of 12,332 million pounds U3O8 (4.7 million t U) are reported from a wide variety of countries and types of uranium deposits. This RAR suggests that there are significant global uranium resources potentially producible at a price of $50 per pound U3O8 ($130 per kg U). Table 1 shows the distribution of global uranium resources. This table, prepared from historic data maintained by the World Nuclear Association, does not include some areas of the world that are currently considered to contain significant resources of uranium such as Mongolia. Table 1 suggests that the United States contains the fourth largest uranium resource in the world with 890 million pounds U3O8, behind Australia, Kazakhstan and Canada in resources that are producible at a price of $50 per pound U3O8 ($130 per kg U). Within the US, the states that demonstrate the greatest amount of uranium resource are Wyoming, New Mexico, Texas and the Four Corners Area of Arizona, Colorado and Utah.

Jan 1, 2007

By J. C. Fox, V. L. Ketellapper

The Summitville Mine, located in the mountains of Southwest Colorado, is an abandoned gold mine. From 1984 to 1992, Summitville was mined for gold using an open pit heap leaching technology. During this period, the volume of acid mine drainage released from the mine significantly increased. This was attributed to the development of the mine pits, the waste rock piles, the heap leach pad, and increased discharges from the drainage adit. The release of this acid mine drainage resulted in the destruction of the aquatic ecosystem for more than twenty miles downstream of the mine. Beginning in 1993, plans were implemented to reduce the amount of acid mine drainage released from the mine. The majority of this work has been completed at the mine. This work includes removal of the Cropsy Waste Pile, closure of the Heap Leach Pad, and revegetation of disturbed areas. Significant improvements to water quality downstream of the mine have occurred resulting from mine reclamation. Aquatic life has returned to portions of the previously mining impacted areas. This study discusses the actions taken to reduce the amount of acid mine drainage generated at the Summitville Mine, the costs of these actions and the resulting reduction in metal loadings. Over 250 samples taken at over 10 locations since 1993 form the basis of this evaluation. Copper loading is analyzed as an indicator of the effectiveness of the actions taken.

Jan 1, 2002

By M. E. Gershon

I would like to congratulate Professor Gershon on an idea that is obviously valuable and is certainly new to me. It has, however, to be further developed, in that one can get fallacious results by limiting oneself to positional weights. In the example problem, the positional weight of one block was 8.6 (the 8.0 value in the paper was incorrect) for 36 units in the cone, for a mean block value of 0.239. The positional weight of the second block was 4.4 for 15 units, for a mean block value of 0.293. Surely one would expect the second block to be rated higher. Additionally, the time value of money adjustment cannot be determined by depth, because the time that a block is mined is determined by the mining sequence, not depth. Figure 1 shows a slightly altered version of Professor Gershon’s pit. The overall value of the pit is the same, as are the total values under each of his blocks. The values in the right-hand cone have been reversed top to bottom, and the values in the rest of the pit adjusted to keep the central cone total unchanged. All values have been multiplied by 10. Figure 2 shows the sequence of mining, at three blocks per year, for Professor Gershon’s pit, and Fig. 3, the sequence for a pit under the other block. Table 1 shows the value and discounted value (157. ) of Professor Gershon’s pit, and Table 2 shows similar values for the second pit. It is apparent that the second pit schedule is more valuable than the first.

Jan 12, 1987

By Steve Kral

About 500 mineral-processing professionals from around the world gathered in San Francisco, CA in Oct. 22-27, 1995 for the XIX International Mineral Processing Congress (IMPC). Following the theme “Leading Mineral Processing into the 21st Century Through Advanced Technology,”186 speakers from 28 countries covered research and development activities and innovative processes. Subjects included comminution and simulation control, physical and chemical processing, flotation practices and fundamentals, precious metals processing and mineral waste and the environment.

Jan 1, 1996

By Ü. Atalay, G. Alaeddinoglu, G. Özcengiz, G. Özbayoglu, Z. Özdemir

In this study, bacterial pre-treatment was applied before cyanidation in order to enhance the solubilization of silver from Artvin-Kafkasör ore Direct cyanidation process was found inefficient to recover silver from the ore. Only 40.36% silver recovery was obtained with direct cyanidation while silver recoveries of above 60% were obtained after bacterial treatment. Thiobacillus ferrooxidans was used as the microorganism in pretreatment tests. Effects of various parameters such as pulp density, duration, inoculum size and particle size on silver recovery have been investigated in this study.

Jan 1, 2001