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By G. L. Oldwright, Virgil Miller

"The data presented in this paper were collected at the smelter of the Consolidated Mining & Smelting Co. of Canada Limited, Trail (Tadanac), Brithish Columbia. The authors are greatly indicated to S.G. Blayrock, vice president and general manager; James Buchman, smelter superintendent; and the staff members mentioned in Report of Investigation 3243. The authors are exceedingly grateful to George Murray, assistant smelter superintendent, for his keen interest, suggestions, and help in carrying out this investigation.INTRODUCTIONThe methods of charign the blast furnaces at Toole and Kelloggs were noted in Report of Investigations 2957 and 3094 respectively. Report of Investigations 2957 cited (4 to 9 inclusive) Hofman's Metallurgy of Lead (p. 226) and the work of A.S. Dwight, Thompson and Sicka, the staff of the Engineering and Mining Journal, and L.D. Anderson, N.F. Collins, in The Metallurgy of Lead (Chas. Griffin, London, 1910) also gives (pp. 247 254) the principles that he believes govern the successful charging of the blast furnace."

Dec 1, 1934

The copper and zinc contents of herbarium specimens of 11 of the 12 Australian species of Polycarpaea were determined in order to establish background and anomalous values of potential use in biogeochemical prospecting.Cumulative frequency curves were used to establish appropriate threshold concentrations for six species including the "copper plant"' P spirostylis. Collection localities of Polycarpaea specimens with anomalous copper or copper-and-zinc values are under investigation.

Jan 1, 1978

By L Avalos, F Melo Miranda, J Misle, M Ingwersen

"Los Pelambres, majority owned by the Chilean company Antofagasta Minerals, is one of the leading producers of copper and molybdenum in the country. It is located high in Andes Mountains, over 3000 metres above sea level, 316 km from the nation's capital, Santiago.The maintenance service supplied to Los Pelambres over the past 13 years has evolved to the point that FLSmidth now provides an integrated service crucial to the management of their customer's assets. The contract initially involved specialty mechanical, electrical and instrumentation maintenance and has developed to encompass maintenance planning and plant control systems.The integrated partnership has delivered excellent results in the areas of improved availability, reliability and return on investment. The operation is now amongst the industry leaders in plant availability.The integrated maintenance service has delivered: An increase in plant availability from 93%, in the first year, to 96.5% today, generating approximately $50 million additional income per year, Quality work procedures, Process and equipment improvements, Major shutdown planning involving more than 700 personnel, 6000 parts and over 2 000 discrete activities.The ongoing success of the partnership requires a high level of understanding between both companies. A commitment to maintain a long term sustainable relationship, based on a ""win - win"" culture, is the key to this success. In this regard Los Pelambres has been a pioneer in establishing and developing such relationships.This case study will share the lessons learnt, the measures taken to improve performance and the best practices adopted during the 13 year evolution of the maintenance service partnership.CITATION:Misle, J, Melo Miranda, F, Avalos, L and Ingwersen, M, 2013. Integrated maintenance services at the Los Pelambres Mine in Chile - The lessons learnt and best practices developed from a long-term sustainable relationship, in Proceedings MetPlant 2013 , pp 296-331 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jul 15, 2013

By Edgar M. Dunn

Part I. Determination oF Gases in Smelter Flues. General considerations. Regular methods for carbon monoxide, oxygen, nitrogen, moisture. Sulphur oxides-—Methods of Lunge and Hempel tried, and either modified or discarded, with reasons therefor. Gravimetric method developed at Anaconda. Description of apparatus used. Brief notes on Richter and Hawley methods for the direct cietermination of sulphur trioxide. Carbon dioxide, with complete notes on same. Arsenic trioxide. Table of percentages of flue gas constituents found at a Western smelter. Some final notations as to the value of flue gas analysis, in connection with an intelligent log of furnace operations. Part II. Notes on the Determination oF Dust Losses at the Washoe Reduction Works, Anaconda, Mont. Discussion of problem, with description of flue system. Various places and methods for determinations of velocity. Instruments used. Procedure in making determinations for velocity. Soundings of cross-section for depths of dust. Calculation of velocities and areas. Daily volume of gas, and plant operations during tests. Place for dust determinations. Apparatus and methods of 'sampling. Results obtained, with table of daily amount of dust losses. Analysis of dusts for values, from various sampling stations. Complete analysis of composite sample of stack emanation. Last analysis totaled to show form in which present. Discussion of recovery as a commercial proposition. References on Pitot tube and calculations.

Jan 1, 1914

The Effect of Ionic Strength of Plant Water on Valuable Mineral and Gangue Recovery in a Platinum Bearing Ore from the Merensky Reef

Sep 13, 2010

By L. K. Weaver

The objective of this study is twofold, The first is to present the composition of the offshore petroleum industry in terms of lease bonuses and the value of hydrocarbons produced from Federal offshore tracts. The second is to present costs of producing petroleum (oil and condensate) in the Gulf of Mexico with a financial analysis of an offshore operation. To prepare a financial analysis of an offshore operation, a model was derived to show the costs necessary to explore, acquire, develop, produce, and abandon a 5,000-acre block and the estimated income from the sale of the hydrocarbons produced. To establish cost and income guidelines for the model, a group of seven oilfields in the Gulf of Mexico ranging from 7 to 75 miles from shore and in water 20 to 130 feet deep was selected for study. A net profit or loss statement was prepared for each field and a discounted cash flow rate of return was calculated. Discounted cash flow rates of return for the seven fields range from 1.1 to 19.5 percent. Under the two specified conditions set up for the model (a 20-year and a 30-year oil depletion model), the rates of return are 17.2 and 13.7 percent, respectively. To January 1, 1959, the top four operators in the Federal offshore areas had paid approximately 40 percent of the amount spent on all lease bonuses and produced about 56 percent of the total hydrocarbons.

Jan 1, 1972

By Donald W. Baggs

This report describes economic mining feasibility studies conducted by the Bureau of Mines to estimate mining costs in the Porcupine mining area of the Juneau Mining District, southeastern Alaska. Investigations were conducted on the economics of three mineral deposit types: 1) massive sulfide deposits containing zinc, copper, and silver, 2) vein gold, and 3)disseminated gold. Cost estimates were developed using the Bureau of mines Cost Estimating S,-stem. Discounted cash flow rates of return were calculated using the Bureau of Mines MINSIM (mining simulation) computer program. Three sets of metal prices were used in the study based on recent historical price information. Underground longhole stope mining of a massive sulfide deposit containing 25 pct zinc. 1 pct copper, and 63.6 g/mt silver was found to be economic at high metal prices at a quantity of 13,200,000 mt ore. Cut-and-fill mining of a 26,400,000 mt massive sulfide deposit was found to be economic at medium metal prices. Open pit mining of a similar deposit was found to he economic at medium metal prices and 26,400,000 mt ore, or at high metal prices and 6,600,000 mt ore. Underground mining of a vein gold deposit was found to be economic at a grade of 17.1 g/mt medium metal prices and 1,320,000 mt ore. Open pit mining of a disseminated gold deposit was found to be economic at a grade of 1.71 g/mt medium metal prices and a quantity of 39,600,000 mt ore.

Jan 1, 1986

By Carmel Bofinger, Maureen E. Hassall, David Cliff, Eric Stemn

e Springer Nature remains neutral with regard to jurisdictional claims in published maps and institutional affiliations.

By R. H. Thurston

Jan 1, 1879

By B. M. Larsen

Jan 1, 1965

By Charles H. White

METHODS in colorimetry are based on the assumption that the intensity of the color of a definite volume of solution is directly proportional to the quantity of the color-producing substance' present. In the preparation and examination of colored solutions there are three ways in which the depth of color may be varied. These variables are; (1) the quantity of the coloring matter, (2) the quantity of the solvent, or the dilution, and (3) the thickness or depth of that portion of the solution examined. With these three variables as a basis, three general methods in colorimetry have been devised, all of which have been used in the determination of carbon in steel. In the first method, the quantity of the solution is kept the same for both the unknown steel and the standard, and the sections compared are of equal thickness, while the quantity of the standard is varied until the color obtained is of the same intensity as that produced by a definite quantity of the unknown steel. In laboratories where this method is used there is pre¬pared and kept for permanent use a series of solutions with varying amounts of the standard, representing the percentages of carbon from the lowest to the highest demanded in that laboratory. These standard solutions are kept in bottles, or test-tubes of equal diameter, and are arranged in series from the faintest to the most intense color, so that it is a simple matter to find the place of the unknown steel in the series, when the solution is of a definite volume and is contained in a tube of the same diameter as the standard tubes. In the second method, equal quantities of the standard and of the unknown steel are dissolved, the standard is diluted to a definite volume, and the solution of the unknown steel is di-

Sep 1, 1906

By C. H. Herty

AN earlier paper on absorption of sulfur by the slag in the basic open-hearth furnace included a brief discussion of the absorption of sulfur during the melting period. The data available at that time showed that the amount of sulfur absorbed by the melting scrap was proportional to the sulfur content of the furnace gases, and the following statement was made: "The conclusion may be drawn from these results that if a gas entirely free from sulfur were used during the melting, any iron in direct contact with the gases would be completely, or nearly completely, desulfurized before it reached the bath proper." The experimental work discussed herein deals with absorption of sulfur by the melting scrap when fuels with both high and low-sulfur content were used. Before taking up the experimental work let us consider the process as a whole as to the "flow" of sulfur. The input of sulfur occurs in: 1, the hot metal charged; 2, the scrap; 3, the stone or lime charged; 4, the ores charged; and 5, the gas. Sulfur goes out of the furnace in the steel, the slag and the gas. The distribution of sulfur in the furnace depends on the sulfur concentration in each of these materials and on the equilibrium conditions between them.

Jan 1, 1926

By B. C. Alberts

Planning the use of South Africa's energy resources, especially coal, is of the utmost importance. The State and the mining industry require reliable official estimates of the economically mineable coal reserves for planning purposes. An overview of al the previous reserve estimates is given, with specific attention to the two most recent estimates-those of Petrick and De Jager. The differences between these two estimates are highlighted. The following important aspects are discussed: . the need to update the information regarding South Africa's mineable coal reserves, . the necessity of understanding and agreement between the coal-mining industry and the State department responsible for updating the coal-reserve data, and . the importance of the specific assumptions used in the updating of the data. A downward adjustment of mineable reserve figures is proposed, and the factors that influence the meaningful exploitation of the coal reserves are discussed. A recommendation is made as to how the Waterberg coal reserve should be exploited, and it is concluded that the Soutpansberg Coalfield cannot be exploited economically in the foreseeable future. An evaluation of the status of the coal-export trade is made and compared with the assumptions and targets set at the start of the coal-exportation scheme. The comparison shows the extent to which the exploitation of the coal reserves is being negatively influenced by the current market situation. Recommendations are made regarding planning for the coal-export market, and the aspects that have to be taken into account to keep South Africa in the export market are identified. Reference is made to the fact that the present unrestrained wage and salary adjustments, which show a total disregard of South Africa's poor productivity record, require serious thought and speedy rectification if the country is to survive in the export market. A salary and wage strategy must be developed to suit the specific needs of the Republic of South Africa.

Jan 1, 1987

By S. P. Kinney

During 1925, at Holt, Ala., the United States Bureau of Mines made a study of the composition of the gases at four elevations in the shaft of a furnace producing foundry iron. A report3 of the work was published in 1927. Some of the results of this investigation are shown in Figure 1. As indicated, the gas at the center of the shaft is low in CO2 content, while the gas in the vicinity of the wall is higher in CO2. The composition of the gases at the top and at the center of the stock column was not far different from the composition of the gas at the top of the bosh. The investigation indicated clearly that lean gas was channeling through the center of the column. More recently the study was enlarged in scope to include investigation of a furnace operating on Lake Superior ores in the Chicago district. A series of planes between the stock line and tuyères was surveyed. Means and apparatus were developed for obtaining, between the inwall and the center of the furnace, the gas composition, gas velocity, and temperature on each plane. A complete report4 of the investigation has recently been published.

Jan 1, 1929

By Guillermo V. Borquez

INTRODUCTION The basic factors that influence the selection of a mining method are the geometry of the ore deposit and the strength of the ore and surrounding rock. These factors usually determine the ore recovery and dilution of the deposit under analysis. For the purpose of this paper we will consider a generalized case of a block caving system and the sequence of the analysis needed to optimize ore recovery. The analysis presented is intended for the preliminary planning stage. The critical factors involved in the caving process are identified and the methods for evaluating the suitability of a particular deposit to the caving method of mining are outlined. The mine design is approached by identifying the two most important, dimensions: the size and spacing of the draw points. These dimensions are estimated from a simulation model which takes into account the ore recovery, the incremental value of the ore recovered, and the safety conditions of the operations. The model estimates the number of draw points required to extract efficiently and safely the broken ore from the caving area. This number is found in the range where the net incremental value of the ore is maximized and the probability of pillar failure does not exceed the limit which is acceptable for a safe mining operation. A mining method and its design are in a continuous process of optimization. This process is unique for each deposit because of the many parameters involved; therefore, a back analysis of the operations should be performed when possible. These analyses will indicate if the ore body responds in a predictable manner to the parameters contained in the geological, fragmentation, and draw point models. SELECTION OF ALTERNATIVE MINING METHODS The selection of a mining method best suited to define the minable ore reserves of a given deposit involves consideration of the geometry of the ore deposit and the strength of the ore and surrounding rock. These two factors are usually sufficient to determine the ore recovery and dilution of the deposit. After analyzing the geological reserves and defining the limits of the deposit, either an open pit or an underground mining method can be selected to define the minable ore reserves. The sequence of steps to follow in the selection of a mining method is presented in Figure I. From an economic point of view, the strip- ping ratio is the determining factor for the selection between an underground or open pit method. In turn, the stripping ratio is a function of the grade distribution and con- figuration of the ore body. In order to obtain a preliminary estimation of the stripping ratio, a parametric analysis could be applied, such as the approach used by Professor E. P. Pfleider (1972) and summarized in Table 1.

Jan 1, 1981

By L. H. ADAMS, C. N. HASKINS, J. K. CLEMENT

In the course of its investigations of the fuel resources in the United States and of the methods by which these resources could be utilized with greatest efficiency, the United States Geological Survey tested a great variety of coals and lignites as gas-producer fuel. Early in these testa it was found that many factors controlling the formation of gas in the generating chamber of the producer and, consequently, having a direct bearing on the management of producer-gas plants, were not as well understood as they should be. The Geological Survey therefore took up a detailed investigation of the chemical and physical processes that take place in the producer, keeping in view not only the possibility of increasing the efficiency of the producer as a source of energy, and the ensuing benefits to the public of cheaper power and greater utilization of low-grade fuels, but also the application of the results to the problems of boiler-furnace operations. - The Bureau of Mines, to which the testing and analyzing of fuels as carried on by the United States Geological Survey has been transferred, is continuing producer-gas investigations at the testing station at Pittsburg, Pa. Results of the gas-producer tests a made at the coal-testing plant erected at St. Louis, Mo., and of a study of some of the problems° that came up in the tests have been published by the Geological Survey. Results of the tests made at Norfolk, Va., and Pittsburg, Pa., and of further studies of particular problems, will be published by the Bureau of Mines. In the gas producer solid fuel is transformed into more readily combustible gaseous fuel. The transformation is relatively slow and consists of several processes: 1. The distillation of the volatile hydrocarbons from the freshly fired fuel at relatively low temperatures. 2. The combustion of the fuel by its combining with the oxygen of the air forming carbon dioxide (COO. 3. The formation of carbon monoxide (CO) and hydrogen (11 3), the essential constituents of producer gas, in accordance with the equations:

Jan 1, 1911

By R. D. West

Miscible displacenzent recovers all oil in the area contacted by the injected .fluid, whereas water or immiscible gas drives usually leave substantial amounts of oil as residual. However, the Door mobility ratios associated with a gas-driven miscible displacement cause the sweep pattern efficiency to be much lower than that obtained with water flooding. One way in which the sweep eficiency in a miscible displacement process can be increased is by decreasing the mobility behind the flooding front. This can be achieved by injecting water along with the gas which drives the miscible slug. This water reduces the relative permeability to gas in this area and thus lowers the total mobility. The main operating conditions for the simultaneous injection -vrocess are that a zone of gas exists between the miscible slug and the leading edge of the water and that a su,@cient amount of gas be injected with the water to form the pas volume which is being left in the water zone. Laboratory model studies have shown that the ultimate sweep pattern efficiency can be as high as 90 per cent for a five-spot flooding system. If gar alone is used as the driving medium an ultimate sweep-out efficiency of about 60 per cent would be obtained in the same system. I INTRODUCTION The miscible displacement processes are a step towards total oil recovery. Conventional gas or water drives usually leave 25 to 5.0 per cent of the oil as residual in the swept portion of the reservoir. This residual can be eliminated if the oil is driven by a fluid with which it is miscible. At some reservoir conditions natural gas will become miscible with the oil. This is the "high pressure gas process".' More often, the oil does not contain enough light hydrocarbons to cause the gas to become miscible with the oil at reasonable pressures. In these cases a small band of fluid which is miscible both with the oil and gas must be kept between them2. Less than 2 per cent of the reservoir volume of the slug material is needed to keep the displacement miscible. Both processes work in the same manner, recovering all of the oil in the portion of the reservoir contacted by the injected fluids. The only difference is the manner in which the miscibility between the oil and the injected gas is obtained. Previous publications have contained detailed descriptions of these processes.1,2,3,4 However, total displacement of the oil in the swept region does not guarantee an efficient recovery process. The amount of oil to be recovered is also determined by the fraction of the reservoir contacted by the flood. This fraction is largely determined by the mobilities of the fluids. (The fluid mobility is the permeability of the rock to that fluid divided by the fluid's viscosity, k/p). This. dependence of the fraction swept on the mobility ratio has been shown in previous studies. Fig. 1 shows the ultimate fraction swept in a five-spot system as a function of the mobility ratio. The small drawings show the location of the areas left unswept for two different mobility ratios. The ultimate fraction of the reservoir swept is here considered to be attained when the producing stream contains less than 5 per cent oil at reservoir conditions. THE GAS-DRIVEN MISCIBLE DISPLACEMENT Since there is no oil left in the swept region after miscible displacement the mobility in this region is very high. It is often 50 times the mobility in the unswept regions. This means that the fraction of the reservoir contacted by the injected fluid will be less for a gas-driven miscible displacement than for a conventional water or gas drive. For a five-spot injection system, water would contact the entire reservoir volume, and the low pressure gas would contact about 90 per cent of this volume, while a gas-driven miscible displacement would only contact about 65 per cent of the reservoir. This poor sweep efficiency often offsets the benefits obtained through miscible displacement. Fig. 2 shows what the recovery curves for the three processes might look like for a five-spot system. The curves show the fraction of the in-place oil recovered as a function of reser-

By William H. Hallahan

The Friedensville zinc mine of The New Jersey Zinc Company is located about four miles south of Bethlehem, Pennsylvania in the Saucon Valley, an infolded and down faulted block of Cambro-Ordovician carbonate sediments surrounded by hills of pre-Cambrian gneiss and Cambrian quartzite except where breached by Saucon Creek at its entrance into the Lehigh River. The Beekmantown formation of lower Ordovician age is the host for zinc and limonite ores, the former occurring in the lower dolomitic portion thereof and the latter in the upper interbedded dolomite and limestone portion of the section. The Beekmantown is overlain unconformably by the Jacksonburg of middle Ordovician age and according to Miller (2) underlain conformably by about 2500 feet of Cambrian dolomites and 200 feet of the basal Cambrian Hardyston quartzite. Igneous rocks, younger than the Cambro-Ordovician carbonate rocks, are absent in the vicinity of the ore deposits. The major structure is a N60°E-trending saddle-shaped doubly plunging anticline overturned to the north. A syncline lies to the north of the anticline. A thrust fault of substantial displacement repeats the south limb of the anticline. The ore consists of sphalerite and pyrite, along with gangue dolomite and quartz, filling and replacing the sedimentary type matrix of a strata-bound solution collapse breccia in the dolomitic portion of the lower Beekmantown. This breccia was produced in pre-Jacksonburg time as a consequence of uplift, erosion, and development of a karst topography and subsurface drainage system in the Beekmantown. Deposits of both zinc and limonite are present on the vertical north limb, on the crest and on the gently dipping (25°) south limb of the Friedensville anticline; consequently, there is no preferred structural setting for mineralization. Rather this distribution supports the view that the folding is post-ore. Features confirming this view are the development of flow cleavage in the breccia matrix, the deformation of some of the sphalerite, and the fracturing of the pyrite mineralization therein. There is nothing to indicate that the age of mineralization is other than Ordovician and older than Taconic deformation. In the absence of intrusive igneous activity younger than the ore host, the writer suggests that submarine volcanic exhalations associated with the volcanic activity recorded in the eastern eugeosynclinal facies of the Ordovician may be the source of the metal ions.

Jan 1, 1968

By F. Weinberg, E. H. McLaren

Dilute binary alloys have been solidified under controlled thermal conditions, and solute distributions, temperatures during freezing and melting, and the position and morphology of the solid-liquid interfaces have been examined. Binary Zn, Sn, and Pb based alloys were investigated with solute additions of AgU°, AU°, Sb'24, TlZo4, snH3, and Zns5. Approximately 150 cc of metal was used in each ingot. Appreciable solute segregation occurred in most of the ingots, depending on the solute concentrations, distribution coefficients. and freezing conditions of the alloys. Measurements were made of the general solute distribution by sectioning layers of material from the ingot circumference and counting the activity of each layer. Local segregation was detected by autoradiographic techniques. Freezing and melting curves for the same alloys were obtained by precision resistance thermometry. These curves are related to the corresponding measured solute distribution curves and those calculated from theoretical expressions. The position and morphology of solid-liquid interfaces during freezing were determined by both decanting and quenching techniques. The results indicate the manner in which freezing and remelting occurred under the conditions investigated. If an isomorphous dilute binary alloy is solidified or melted under equilibrium conditions, the concentration of solute in both solid and liquid would be uniform at all stages in the freezing process. The solute concentrations during freezing could be determined from the phase diagram of the alloy, and the freezing curves would have an alloy slope defined by the solidus and liquidus points. In practice, however, solidification under equilibrium conditions rarely, if ever, occurs, primarily because of low rates of diffusion in the solid and incomplete mixing in the liquid. As a result, the solute distribution and the shape of the freezing and melting curves obtained in a given system depends on a number of factors. These include, in addition to the extent of diffusion and mixing, the initial solute distribution and concentration, the distribution coefficient, the rate of freezing or melting, supercooling of the liquid, the morphology of the solid-liquid interface, and the effect of annealing. The purpose of the present investigation was to determine, experimentally, actual solute distributions and thermal curves for a series of dilute binary alloys, using tracer techniques to determine the solute distributions and precision resistance thermometry for the temperature measurements. In addition, observations were made of the positions and morphologies of the solid-liquid interfaces during freezing and remelting. The distribution of solute in a dilute alloy, assuming complete mixing in the liquid, and negligible diffusion in the solid, is given by1 where Co is the mean solute concentration of the alloy, C, is the solute concentration in the solidg is the fraction of the ingot solidified, and ko is the equilibrium solute distribution coefficient (assumed to be independent of concentration for dilute alloys). For the case of partial mixing, ko can be replaced by an effective distribution coefficient k. If solute transport in the liquid is assumed to be entirely due to diffusion, with negligible diffusion in the solid, then the solute distribution in the solid is given by Tiller et a1.' as where R is the rate of advance of the solid-liquid interface, D is the solute diffusion coefficient in the liquid, and x is the position of the interface from the start of freezing. The temperature at the interface TI is related to the solute concentration of the solid at the interface C, by the expression where Tm is the freezing temperature of the pure solvent and ml is the slope of the liquidus for the binary alloy being considered. Accordingly, by relating TL to the actual temperatures measured in the present experiments, solute distributions can be deduced from the temperature measurements. Using the above expressions, the measured solute distribution of a given alloy may be compared to the theoretical distributions, as well as to the distribution calculated from the observed freezing curve. EXPERIMENTAL The solvent materials used in making up the binary alloys were C. P. New Jersey zinc (99.999 pct), Vulcan Extra Pure Tin (99.999 pet), and Tadanac lead (99.998 pct). The solute elements were of 99.99 pct purity or better, and were activated in the

Jan 1, 1963

By G. D. Naugle, L. C. Atkinson

Introduction Deep open-pit mines invariably affect the local and regional hydrologic systems. Pit dewatering, occurring during mining operations, puts an obvious hydrologic stress on these hydrologic systems. However, post-mining hydrologic impacts resulting from the pit refilling with groundwater following the cessation of mining activity can also be significant. The prediction of the rate at which the post-mining pit will fill with groundwater is a critical aspect of assessing the long-term hydrologic impacts. Numerical groundwater flow modeling provides a method for predicting the groundwater refilling rate of the pit. The rate at which pit "lakes" fill depends on several factors: •the rate and duration of pit dewatering; •the depth and size of the ultimate pit and •the pre-mining hydrologic regime. These factors can be incorporated into a detailed numerical groundwater flow model that can then be used to assess the effects of dewatering and post-mining recovery on the local and regional hydrologic systems. A sufficiently detailed, numerical groundwater model provides the oportunity to: •account for complex geology near the pit; •assess the impact of active pit dewatering and •predict the long-term impacts of post-mining groundwater flow into the pit. A detailed groundwater model incorporating these items has been developed and applied at an operating open-pit mine. Developed by Durbin and O'Brien (1987), the three-dimensional, finite-element, groundwater flow model was used to represent the hydrologic system of an approximately 253-km2 (98-sq mile) area surrounding the pit. Historical groundwater elevation data, stream flows and meteorologic, geologic and geophysical data were used to establish the dimensions and initial conditions for the model. Steady-state conditions, representing the pre-mining local and regional hydrologic systems, were simulated using the initial conditions incorporated into the groundwater model. The groundwater model was then utilized to simulate various dewatering programs, to predict the filling rate and the groundwater depth in the ultimate pit once mining activities are complete and to assess the long-term impacts on the regional groundwater flow system. Development of pit lake model Groundwater modeling efforts were completed in two phases. The first focused on pit dewatering activities, while the second phase concentrated on the post-mining effects on the hydrologic system. The final estimates of groundwater elevations calculated during the pit dewatering simulations were used in predicting the post-mining recovery of the hydrologic system. The groundwater model was also modified prior to the second phase to account for the volume of rock removed during mining activities. To account for the actual volume of rock mined, the geometry of the post-dewatering model grid was modified to approximate the final pit geometry. The depth and width of the ultimate pit were divided into eight idealized stages that represented significant changes in the bench geometries. These eight stages were then introduced sequentially into the model according to the predicted water elevations within the pit. In this way, changes in the volume and depth of water within the pit were accounted for through time. Once the ultimate pit geometry was accounted for in the model, it was necessary to assign new hydraulic characteristics to those parts of the model grid (elements) that represented excavated rock. The solution of the numerical model requires that finite hydraulic conductivity values be assigned to the portion of the groundwater model that represents excavated rock. Therefore, the calculated groundwater elevations differ, somewhat, between the edges and the center of the open pit. These model-calculated water elevations at the edge and in the middle of the open pit represent the elevation of water that would occur in the pit lake. To minimize the error in the estimated level of water within the pit lake, the hydraulic conductivity was increased to a value that would: •minimize the predicted difference between the groundwater elevations across the open pit and •produce a numerically stable solution. Specific storage is the hydrologic parameter that accounts for the water produced by compaction of the aquifer matrix. To predict the groundwater volume that would flow into the ultimate pit, this parameter was assigned a value equivalent to the compressibility of water. This value of specific storage reflects the post-mining groundwater storage occurring as an open body of water. Additionally, a specific yield of 1.0 was assigned to the pit elements to represent the 100% porosity of the open pit. In

Jan 1, 1994