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By Martin B. Zimmerman

It is becoming evident that traditional methods of econometric analysis are inadequate in estimating-the long-run supply of minerals. Econometrics is limited to extracting estimates from observations of the theory predicts that the least cost deposits will be exploited first.1 The estimation of parameters from past data then must lead to biased projections of the future. Mining conditions in the future will be worse than in the past and it is the impact of the future conditions on cost that we need to know. In this paper we present a methodology for dealing with this problem. We estimate a "cumulative cost" function that yields the long-run marginal cost of mining as a function of cumulative output produced over time.2 This is relatively easy to do for coal mining in the United States where exploration for new deposits is not a major factor in supply. The first step of the procedure relates cost to the rate of output and to the geological conditions of the mine. The second step combines geological information on remaining deposits with the cost function to yield the cumulative cost function. A central theme of both these steps is the interaction of economics with geology. The best deposits are selected first and in each step of the estimation allowance must be made for this fact. The problem of depletion is part of the general problem of structural change. The change through depletion in the nature of the exploitable mineral stock is only one factor that calls for an analysis sensitive to the changing circumstances of the industry. The factors impinging most heavily on the coal industry today are the strict environmental regulations legislated

Jan 1, 1977

By K. Burke, K. L. Myers, R. Espell

Barrick's Goldstrike AA heap leach pad is an HDPE-lined facility that was used for gold recovery from low-grade ores using conventional cyanide heap-leach technology. The pad was operated from 1987 through 1999. Goldstrike prepared a unique reclamation plan for the AA leach pad facility to provide for a cost-effective, environmentally sound facility closure and to fulfill the State of Nevada's closure requirements. The traditional Nevada leach pad closure procedure involves rinsing to reduce the cyanide concentration to an acceptable level followed by reclamation. However, rinsing does not typically reduce certain metals and compounds, and some constituents may increase in concentration with rinsing. A closure scenario to effect limited infiltration through installation of a capillary cover for the leach pad was developed. The capillary cover was constructed by placing a fine-grained layer above the coarse-grained spent leach material to create a capillary break. In an arid climate such as in Nevada, the fine-grained layer will rarely reach saturation, and the cover will be able to store water during precipitation events and will allow removal of water through evaporation and transpiration (ET) during the growing season. A 1.22-m (4 ft) cover thickness was selected as the optimum design cover thickness. The drainage network on the surface of the recontoured AA heap leach pad was specifically designed to enhance the re vegetation, reclamation and sediment yield control aspects of the AA heap leach pad closure, while resulting in a natural looking surface that is compatible with the surrounding topography. Adjacent natural topographic features were used to establish target design parameters for recontouring and erosion-control design.

Jan 1, 2002

By G. S. S. Raman, M. S. Klima

The effects of pressure and cake thickness were studied using a split-plot design for pressure filtration of a fine refuse slurry obtained from a bituminous coal preparation facility. Pressure filtration offers the potential to significantly reduce the moisture content of the fine refuse stream with the addition of little to no flocculants, thereby minimizing the volume of refuse stored in impoundments and providing a recycle water stream for use within the plant. Experimental results showed that the final cake moisture decreased with increase in pressure and decrease in cake thickness. In addition, specific cake resistances calculated for the various tests indicated an increasing trend with increasing pressure and decreasing cake thickness. With increase in cake thickness, an increase in cake moisture and the weight of solids deposited in the cake was observed. An ANOVA analysis indicated that both cake thickness and pressure significantly affected the cake moisture, while only cake thickness statistically influenced the weight of solids in the cake. A regression model was developed with an R2 value of about 0.99 to relate cake thickness, the single largest statistically significant factor, with the weight of solids in the cake. Filter capacities were calculated in terms of the solids deposited per unit time based on a product specification of 25 percent cake moisture, and it was observed that the thinnest cake had the highest filter capacity.

By D. E. Healy

In 1989 an Autogenous Grinding Task Force was formed to study potential autogenous grinding circuits to replace the existing semi-autogenous grinding at the National Steel Pellet Plant located in Keewatin, Minnesota. Pilot plant test work performed by the Hanna Research Center (HRC) indicated that fully autogenous grinding with pebble extraction and crushing could feasibly be incorporated into the existing circuits. In 1990, one of the existing ten grinding mills was converted for an in-plant test incorporating pebble crushing. The full scale test work conducted in 1990 and 1991 verified the pilot plant projections with primary mill throughputs from the fully autogenous circuit utilizing pebble crushing being greater than throughput of the existing SAG mill circuits.

Jan 1, 1992

By Gary Peach, Peter Mason

Planning and estimating progress rates and costs for tunnelling is always best done based on experience and historic precedent. Much precedent is available but the use of data monitoring, data mining and other retrieval techniques on modern TBM machines can enable this to be further expanded and refined. Advance rates can be related much more closely to rock types as can associated rock bolt and mesh/ shotcrete requirements. This in turn can lead to a further refinement in best likely combinations of equipment, such as shotcrete jumbos, better selected for the anticipated conditions, rather than the potential advance rates of the machine. Efficient and timely construction of the tunnel will also be linked to times needed for finishing work, such as contact and void grouting and to how this can be initiated while the TBM is still in operation, in order to save overall construction time. Indeed, this in turn may affect the most preferable layout and construction details of the permanent tunnel. Considerable experience has been gained in the use of two large diameter TBM tunnel drives through the Himalayas at the Neelum-Jhelum hydropower project in Pakistan. All the above aspects will be discussed in the paper based on the considerable experience and data which came from that operation, including appropriate measures in highly stressed areas potentially subject to rock burst.

Jun 13, 2021

By W. Collingwood

As openpit mines continue to grow deeper and productivity continues to increase, the management of respirable dust and noxious gases can become a challenge. The natural wind flow in openpit mines is often recirculatory. This recirculation traps dust and gases in deep openpit mines. This problem is particularly acute in high-latitude openpit mines, due to a phenomenon known as atmospheric inversion. In the winter, arctic and subarctic regions experience short days and long nights. This deficit of solar radiation creates a situation in which cold, still air accumulates near the bottom of the pit. In openpit mines, this can result in unhealthy accumulations of gases and other contaminants. In this study, the natural wind flow patterns in and around openpit mines were investigated using a computational fluid dynamics (CfD) program. Two-dimensional models were created for a variety of pit geometries. The effect of depth, slope angle and wind speed on the recirculation pattern in the pit was examined.

Jan 1, 2012

By Shiwei Wang, Ziwen Li, Chenhu Zhang, Peng Chen, Chengyong Wang

Accurate prediction of the partition coefficients in coal gravity separation is key to improving the preparation efficiency of coal and promoting its clean utilization. In China, the approximate formula method (AFM) has been applied to the calculation of coal gravity separation partition coefficients for decades, but its accuracy is limited. To improve the accuracy of the predicted partition coefficients, and thus the product yield of coal preparation plants, the present study proposes a modified hyperbolic tangent model to derive direct calculation formulas for the heavy medium and water medium separation partition coefficients. Four groups of coal samples were used to verify the accuracy of the direct formulas, and their predictions were compared with those of the AFM. The results showed that the partition coefficients and partition curves predicted by the direct formulas were closer to the actual values than those predicted by the AFM. Further, when the cleaned coal ash content was basically the same, the cleaned coal yield predicted by the direct formula was lower than that predicted by the AFM, and was also closer to the actual value. These results demonstrate that the direct formula produces more accurate results compared to the conventional AFM. As this formula has a reasonable theoretical basis, is simple to apply, and produces accurate results, it is expected to help improve the design and management level of coal preparation plants.

Oct 15, 2022

By Simon E. Rodriguez

This paper presents sulfur isotope data obtained for the deposits and country rocks of the Bailadores and Aroa Mining District, Venezuela. The study is a part of an extensive geochemical investigation of the mineralized belts that has as their aim the elucidation of the genesis of ores. The Bailadores deposits are located at about 600 Km. west-southwest of Caracas, or about 10 Km. southeast of the Town of Bailadores, Merida State (Fig. 1). The mine concession is held by the Ministry of Mines and Hydrocarbons (De Lima I, De Lima II and La Cascade areas). In a straight distance the mine is about 90 Km. southwest of Merida, the capital of Mérida State. From Mérida to Bailadores, through La Victoria and Tovar, it takes 3 hour by car. From Bailadores to the mine site it is about 40 minutes by jeep. The Aroa deposits are located at 4 Km. east of Aroa Village, Yaracuy State, about 220 Km. west of Caracas. It is 83 Km. from San Felipe, the capital of Yaracuy State to the mine (Fig.2). The mine is about 600 m above sea level, situated on the northern foot of the E-W trending Andes. The topography of the mine are shows a mature stage and is somewhat rugged. As the climate is warm and humid, the area is thicky covered with jungle.

Jan 1, 1974

By Herbert F. Nyberg

The Black River Falls Mine is owned and operated by Jackson County Iron Company, a wholly owned subsidiary of Inland Steel Company. It is located in vest central Wisconsin, seven miles east of the city of Black River Falls. This is the first taconite plant in the state of Wisconsin. Iron mining started in this area in 1856 but soon failed due to the low grade ore and crude methods of mining and transportation. Inland Steel Company acquired the property in 1939. Some drilling and test work was done in 1940, but the economics of developing the property at that time were not favolable. Improved methods of concentrating and pelletizing in the 1960's made it possible to build the present plant. Due to the limited ore reserve of 45,000,000 tons with expected recovery of 33% it was necessary to limit the capacity of the operation to 750,000 TPY with a life of 20 years. There was some concern about an operation of such limited capacity. A million ton operation was considered the minimum size plant for a profitable operation.

Jan 1, 1976

By J. Bilant, B. Carlson, N. Gow

High Pressure Grinding Roll (HPGR) technology has been evaluated to replace tertiary cone crushing at Coeur Rochester operations (Coeur Rochester, Inc.) as an opportunity to improve silver recovery. In 2016 and 2017, onsite and third-party pilot programs were initiated to evaluate the impact of metallurgical recovery and heap leach performance of crushed product generated by an HPGR. In 2018, Forte Dynamics evaluated the HPGR laboratory metallurgical data from Coeur Rochester and all third parties to develop field-scale leach curves based on new crushing scenarios. The laboratory recovery test work data was consolidated and scaled to field conditions to determine projected heap leach performance. Based on the metallurgical testing conducted, the replacement of current tertiary cone crushers with HPGR technology has the potential to increase Coeur Rochester’s silver recovery rate by 8%, from 61% to 69%, and shorten the overall timing of silver recoveries from 20 years to just over two years. INTRODUCTION The Rochester mine, and associated valley fill heap leach facilities is an open-pit silver and gold mine located in Pershing County, Nevada, approximately 26 miles northeast of the city of Lovelock. In 2017, Rochester produced approximately 4.7M troy oz silver (Ag) and 51k troy oz gold (Au). Ore mined at Rochester is either fed directly to the primary X-Pit crusher, crushed at an in-pit crusher system, or placed directly onto a heap leach pad as run-of-mine (ROM). Both crushing circuits are comprised of three stages of crushing to produce a nominal 3/8” ore. Primary crushing consists of jaw crushing while secondary and tertiary crushing are composed of varying cone crushers. Historical recovery rates of crushed Rochester ore average 61.4% Ag and 95.9% Au over a period of 20 years [1]. These rates were derived from third party verification of in-house metallurgical test results, crusher production and heap leach production results from 1986 through 2004 in a private report for Rochester produced by KD Engineering Co., Inc. Recovery rates are regularly verified by compositing daily samples of ore delivered to each leach pad from the X-pit crusher, in-pit crusher, and/or ROM. A proportionate split of each sample is used to generate monthly composites for metallurgical analysis utilizing column leaching and bottle roll leaching. These tests are run in a manner analogous with production conditions and deliver test results that provide expected heap leach performance. As the Rochester oxide deposit has been mineralogically and metallurgically consistent, the gold and silver recovery trends have also been consistent over the life of the project. In 2016 and 2017, onsite and third-party pilot programs were initiated to evaluate the impact of HPGR technology to metallurgical recovery and heap leach performance of crushed product in an open circuit configuration, as a potential replacement for the currently in place tertiary cone crushers. During these years, HPGR was evaluated in three phases of test work.

Jan 1, 2019

By Christopher J. Bise

It is somewhat reassuring to note that the United States has an abundance of coal reserves which can be used to bridge the widening gap between the nation's energy production and consumption. The U. S. Bureau of Mines has recently established a demonstrated coal reserve base of 434 billion tons. (31)(Figure 1) As can be observed, the major percentage of the reserves is in the West. Although surface coal mining in the West has been gaining increased attention in recent years, western reserves that must be deep mined are significantly greater. Though deep mining activity in the west is very limited and accounts for less than 2% of the annual production, the primary reason, quite probably, is the greater thickness of the seams as compared to those found in the East. Forty-five billion tons of the western demonstrated coal reserve base for underground mining are in seams greater than 10 ft. in thickness. (37) Thick coal seams do not lend themselves readily to extraction with the underground mining methods practiced in the United States today. Experiences from the practices in other countries can serve only as guidelines. Consequently, a concerted effort is necessary to overcome the technical difficulties.

Jan 1, 1975

By Edward S. Hsia

Pluor Utah has been closely associated with the engineering and construction of earth bottom thickeners for porphyry copper operations, for the counter-current decantation circuit of a gold cyanide plant and similar applications in mineral processing. During the past 13 years, some 18 of these thickeners have entered service and the bottoms have proved to be watertight structures with no particular problems encountered. In the course of this work, technological advances have been realized in the procedures for engineering and construction of these units to assure watertight basins and achieve all possible economies in construction. This paper describes the successful application of engineering and construction techniques for a number of these thickeners.

Jan 1, 1977

By Paul M. Jr. Musgrove, Thomas D. Henderson, Mark B. Nesbitt

Introduction History. Vat leaching operations began in October 1953. Since that time, about 4.2 million tpy of oxide ore averaging 0.72% copper have been treated. Economic conditions, improved operating methods, and increased production facilities have allowed the ore grade to be reduced over the years. The average ore grade was 0.84% Cu during the 1950s, 0.68% in the 1960s, and was 0.50%, or below, during the 1970s. By 1961 some 35 million tons of oxide ore had been removed from the mine, exposing the underlying sulfide ore body. A 5,000¬ tpd concentrator was constructed and put into operation to treat the sulfide ore in November 1961.1 The original crushing plants were used for both oxide and sulfide ores, but by adding a third crushing shift, six days per week, the total ore treated was raised from 12,000¬17,000 tpd. Dwindling oxide ore reserves prompted expansion of the concen¬trator to 13,500 tpd in 19672 so that exhaustion of known reserves of both ores would occur at nearly the same time. Again, no additional crushing facilities were added during the expansion. It was anticipated that the treatment rate of oxide ore would be reduced from 12,000 to about 9,000 tpd. Instead, the addition of two crushing shifts per week and improved operating procedures have increased the treatment rate to 26,500 tpd-13,000 tpd to the leaching plant and 13,500 tpd of sulfide ore treated in the concentrator. Geology. The Yerington ore body is a disseminated porphyry copper deposit where a quartz monzonite porphyry intrusive contains the major part of the ore values, with some copper mineralization found in the adjacent granodiorite and other rocks. The geology of the Yerington Mine was aptly described by J. R. Wilson (Mining Congress Journal, June 1963). The ore body is one of few porphyries where geologic conditions were optimum for the formation of an important concentration of oxidized ore with minimum economic effect of secondary enrichment. The principal mineral in the oxidized zone is chrysocolla, which irreg¬ularly occurs dispersed throughout the rock and as narrow seams along fissures. The grade of the ore in the oxidized zone is generally the same as in the primary sulfide below, although some massive concentrations do occur as vein-like fissures and as filling of breccia zones. Cuprite, tenorite, melaconite, malachite, and azurite occur but are not abundant. Lying between the chrysocolla horizon and the primary sulfide is a transition zone whose limits are irregular but seldom with vertical range exceeding 50 ft. Chalcocite, cuprite, melaconite, native copper, and chrysocolla occur superimposed on the primary sulfides. Chalcocite replacement of chalcopyrite, as well as secondary chalcocite, do occur. The primary sulfide minerals are chalcopyrite and pyrite, with minor amounts of bornite and covellite. There are no appreciable gold or silver values, and molybdenite has rarely been noted.

Jan 1, 1985

By E. G. Kelly

This paper describes the performance of a modern compound trough profile spiral that is optimized for fine coal cleaning. Reduced performance (Tromp) curves are presented to compare these new spirals with older designs, as well as with other commonly used coal cleaning equipment. As an aid to equipment selection and circuit design, data are also presented to show how variables, such as flow rate, pulp density, and particle size affect efficiency and the separation specific gravity.

Jan 1, 1984

By L. D. Patts, S. J. Janisko, E. G. Cauda, A. D. Bugarski, G. H. Schnakenberg, J. A. Hummer

Filtration systems with disposable filter elements (DFEs) are used in the underground coal mining industry to control particulate matter emissions from diesel-powered permissible and nonpermissible coal mining equipment. This study was conducted in underground mine conditions to evaluate three types of high-temperature DFEs used in those filtration systems. The DFEs were evaluated for their effects on the concentrations and size distributions of diesel aerosols and concentrations of nitric oxide (NO) and nitrogen dioxide (NO2). Those effects were compared with the effects of a standard muffler. The experimental work was conducted directly in an underground environment using a unique diesel laboratory developed in an underground experimental mine. After an initial DFE degreening period, the filtration system with all three DFEs was found to be very effective at reducing total mass concentrations of aerosols in the mine air. The effectiveness of DFEs in filtering aerosol mass was found to be a function of the engine operating conditions. The efficiency of the new DFEs significantly increased with operating time and buildup of diesel particulate matter in the porous structure of the filter elements. A single laundering process did not exhibit substantial effects on the performance of the DFE elements. The effectiveness of DFEs in removing aerosols by number was strongly influenced by engine operating mode. The concentrations of nucleation mode aerosols in the mine air were found to be substantially higher for DFEs when the engine was operated at high-load modes rather than at low-load modes. Initial heating of certain DFEs resulted in visible white smoke and substantially elevated aerosol number concentrations. The effects of the DFEs on the total concentration of nitrogen oxides (NOX) were found to be minor. The NO2 fraction was found to be generally lower for the DFEs than for the muffler. The engine-out NO2 fraction of the total NOX was found to be substantially higher for low-load modes than for high-load modes..

Jan 1, 2011

By H. S. Castro

The by-product molybdenite technology in Chile is reviewed. Flowsheets, reagents and some operation features for the actual six Chilean moly plants(Chuquicamata, El Teniente, El Salvador, Andina, Las Tortolas and Pelambres), are presented. The chemical and physicochemical factors involved in the depression of copper and iron sulfides during the selective flotation of molybdenite, with sodium sulfide, sodium sulfhydrate, anamol-D and LR-744, are discussed. Advances in technology to solve problems of excess of lime (precipitation with sulfuric acid); excess of residual flotation reagents (activated carbon and kerosene); and optimization of depressant consumption (Eh control, nitrogen and low-oxygen air), are discussed.

Jan 1, 1998

By Matias Pentinnen, Wayne Dickinson

Electrowinning is a key process for copper production contributing roughly twenty percent globally and thirty to forty percent of production in the Americas1. Copper grade is assessed largely on deposit smoothness and on level of impurities in the metal. Smoothing aids are essential to create dense, low porosity deposits and guar has long been the material of choice for this purpose. Guar is, however, poorly stable under the strongly acidic electrowinning conditions, underlying ongoing research to identify an effective, acid-tolerant replacement. Current efficiency is also a critical factor in controlling copper production costs and chemical additives should ideally improve current efficiency as well as improving deposit smoothness.

Feb 1, 2025

By P. E. Damon

An arsenal of isotopic techniques is avail-able that can date the full range of hydrothermal ore deposits. The following are prerequisites for ideal results: (1) Careful collection of samples that are related to the event in a field and process context; (2) Optical examination of the component selected for dating to be sure that it has not been altered; (3) Careful preparation of a pure component of the sample selected for dating; (4) Precise and accurate measurement; (5) Knowledgeable interpretation of the data with reference to decay systematics, properties of the component analyzed as well as the field context and thermochemical environment of the ore process.

Jan 1, 1992

By Roy H. Paddon

This paper describes the 6, 000 tpd, three stage haulage system at the Ozark Lead Company's mine in southeastern Missouri. It details the methods and problems pertinent to the loading, hauling, sorting, crushing, storing and hoisting of waste rock and ores obtained from underground development headings and production areas. The Ozark Lead Company is a wholly owned subsidiary of the Kennecott Copper Corporation. Its mine is located nine miles south of Bunker in Reynolds County at the south end of the "New Lead Belt" mining district. Development started in May, 1964, by drilling a seven foot diameter (six foot cased), 1, 250 foot development-ventilation shaft. Drifting began in August, 1965, to obtain a bulk ore sample, to test ore continuity, and to drive two mine levels while the pre-grouted production shaft was being sunk. Over 7, 900 feet of 12 by 24 foot drifting was done from this small shaft. In March, 1966, a 20 foot diameter, three compartment, concrete lined production shaft was collared. It intersected existing development headings at a depth of 1, 195 feet in April, 1967, and was completed along with a concentrator and other facilities in June, 1968. Mine production began that month, with 5, 000 tpd reached in 1970.\

Jan 1, 1971

"Twin Creeks is a gold mining complex that is part of Newmont’s Nevada operations. The mining complex is composed of two openpits, Mega and Vista, external ore sources and a set of existing stockpiles, all of which provide ore for an autoclave, an oxide mill and a heap leach. Mega pit provides sulfide ore for the autoclave, whereas Vista pit provides oxide ore for the oxide mill and the heap leach. Both pits operate the same mining equipment and their extraction sequence must, therefore, be optimized simultaneously. Stringent blending requirements are associated with the operation of the autoclave for the sulfide ore. Strategic blending optimization on a large scale has brought significant value to the operation by increasing synergies. This paper presents the implementation of a stochastic optimization framework for the long-term production planning at Twin Creeks that simultaneously optimizes mining, stockpiling, blending and processing decision variables. The uncertainty and variability associated with the different sources of material are incorporated in the optimization model by means of stochastic simulations. The stochastic solution generated shows substantial benefits by increasing forecasted recoverable gold, leading to increased cash flows, while reducing the risk of not achieving forecasts by increasing the probabilities of meeting production and blending targets.IntroductionMine operations consist of multiple components such as mineral deposits, stockpiles, external sources, mills and autoclaves. These components represent a mineral value chain, also known in the literature as a mining complex (Montiel and Dimitrakopoulos 2015; Goodfellow and Dimitrakopoulos 2016). The material extracted from the mines flows through the different components and is transformed into sellable products. This flow of material in a mineral value chain involves complex, nonlinear transformations that render the use of industry-standard optimizers difficult. Furthermore, the optimization of a mining complex can be formulated as a mixed-integer, nonlinear optimization program with millions of integer variables representing mining, blending, stockpiling and processing decisions, prohibiting the use of conventional nonlinear optimization approaches. Previous methods that consider the joint optimization of multiple components of the mineral value chain are available (Hoerger, Seymour and Hoffman, 1999; Chanda, 2005; Stone et al., 2005; Wooller, 2005; Whittle, 2007, 2010; Bodon et al., 2010). Although these methods add value by increasing the synergy between the different components of the mining complex, they rely on simplifying assumptions that constrain their efficacy, while ignoring a most critical aspect. The latter aspect includes the geological uncertainty and actual variability of the materials in the mineral deposits mined and are supplied to the related mineral value chain, similarly to conventional life-of-mine planning methods used for individual mines (Dimitrakopoulos, Farrelly and Godoy, 2002; Godoy and Dimitrakopoulos, 2011; Spleit and Dimitrakopoulos, 2017). Additional simplifying assumptions may include: the consideration of a limited number of components, the combination of different methods, the aggregation of variables, fixed production schedules the definition of policies prior to optimization, such as cutoff- grade definition or destination of mining blocks, the linearization of nonlinear transformation functions and/or the disregard for the uncertainty of several critical parameters (Whittle, 2007). The simultaneous optimization of multiple components of a mining complex was introduced by Newmont. Hoerger, Seymour and Hoffman (1999) implemented a mixed-integer programming formulation to optimize Newmont’s Nevada operations. The model is based on the work of Urbaez and Dagdelen (1999) and simultaneously optimizes the extraction of materials from multiple deposits and the delivery of ore to multiple plants, as well as as"

Dec 1, 2018