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By Brian Reid

"This paper studies the progress made by hydraulic shovel manufacturers over the last few years, particularly in the areas of productivity increase and operating cost reductions.Five years ago, it was widely considered that the unit production cost for a hydraulic shovel in a mining application was double that of a cable shovel in the same application. The hydraulic machine manufacturers have taken up this challenge, and modern improvements in design, coupled with changes in maintenance techniques, have narrowed the gap to the point where it is valid to ask the question ""will the cable shovel exist in ten years' time?"".Historical BackgroundThe history of the cable shovel goes back to the early 19th century and the industrial revolution. Early machines were steam-powered and ran on rails.The ""power shovel"" and the ""dragline"" were the only mechanical means of mass excavation until the early 1950s. Rail-mounted units gave way to crawlers, which increased flexibility considerably. Some shovels were redesigned as backhoes, and some became extremely large when designed as stripping shovels primarily used in North American coal mines.In the early 1950s growth in hydraulic technology spawned the introduction of the hydraulic excavator, which soon began to replace the smaller cable shovels and backhoes used in the construction industry. They were completely self-contained with either a gasoline or diesel engine and, being much lighter, were more readily transportable.As the excavators grew in size from its early models of 0.5 m3 or less, many of the cable shovel models were discontinued. This was not surprising for several reasons, but the most obvious is that both were often made by the same manufacturer."

Jan 1, 1988

By C. Pelley, S. Kelebek, G. Kunzel, S. Gregersen, M. Pollanen

A series of small-scale tests was conducted to evaluate the effect of micro-fracture development, imposed by blasting, on the material properties of the rock as well as its resistance to crushing and grinding. The uniaxial compressive strength, tensile strength, Poisson’s ratio, Pwave, S-wave velocity, point load test index, crushing, and grinding work indices were evaluated for samples without any blasting load and for samples which had received a variety of blast loads ranging from a single impact with a lowstrength detonating cord to a combination of impacts at different times and significant impacts with powder factors close to those used in practical blasts. It was shown that the rock mechanics parameters change even at low impacts, suggesting easier crushing of the rock. Damage values appear to be affected by timing as stress wave superposition affects the damage process. Fragmentation is achieved at higher impact levels while work index modification of the fragments is also observed when high powder factors are used, suggesting improvements in grinding.

Jan 1, 2006

By S. Banisi, M. Poorkani, M. Farzaneh, M. Askari

"At the Sarcheshmeh copper mine 40,000 t of ore per day with an average grade of 0.9 % Cu is ground to 70% passing 75 microns. Grinding is effected in 8 parallel lines of 8 x 5 m ball mills in a closed circuit with cyclones. The make-up balls are 80-mm forged alloy steel balls and the average ball consumption is 750 gh of ore ground. In order to optimize ball size distribution inside the mill, based on the previous investigation instead of using only 80 mm make-up balls a combination of 80 and 60 mm (75% to 25% by weight) ball charge was used in one of the mills. To evaluate the performance of three ball types namely forged alloy steel balls (A), ductile cast iron (C), high chromium cast iron (D) and also a combined (80 and 60 mm) charge of forged alloy steel balls (B), four identical parallel mills were charged accordingly. These four mills were sampled during a period of six months. The amount of material finer than 75 microns of cyclone overflows was selected as a criterion for the performance evaluation. It was observed that in a mill with the combined make-up balls the amount of fine produced was 4% higher than the other mills. The consumption of ball types A, C and D were 730, 710 and 534 g/t, respectively. The results of laboratory flotation tests showed that due to an improvement in the production of fines an increase of 1.4% in the overall copper recovery is obtainable. Due to the promising results, the new make-up ball regime was implemented in all mills in the plant.INTRODUCTIONThe Sarcheshmeh copper ore body which may rank as third or fourth largest in the world contains 1 billion tonnes averaging 0.90 % copper and 0.03 % molybdenum. It is located southeast of Iran and has been processing 40,000 t/d of ore since 1982. The mine produces 100,000 tonnes of copper and 2,200 tonnes of molybednite concentrate per year. After three stages of crushing, the ore with a F80 of 12.7 mm is fed to 8 parallel ball mills in closed circuit with cyclones to produce a product 70% finer than 75 microns. The concentrate of the first stage of flotation is reground and the tailing constitutes the main component of the plant final tailings. A concentrate with an average grade of 31% Cu is obtained after cleaning and re-cleaning stages. Total recovery of the plant varies between 83 and 87 % depending on the operating conditions and ore type."

Jan 1, 2002

By John G. Payne

The Pebble deposit contains three contiguous zones. In the Pebble West zone, granodiorite stocks intrude gently warped volcanosedimentary rocks that had been intruded by diorite sills and two felsic intrusion breccias. In the Central zone, volcanosedimentary rocks are cut by diorite and granodiorite sills; the latter are genetically associated with the granodiorite stocks. In these zones, widespread potassic alteration consists of brown biotite and K-feldspar, with abundant quartz-carbonate-sulphide vein stockworks. Chalcopyrite/pyrite ratios range from 2/1 to 1/20. Diorite hosts the highest-grade mineralization, with much of it being disseminated. Patchy, later phyllic alteration is moderate and commonly depleted in copper. The Pebble East zone occupies a granodiorite stock and intruded host rocks below an eastwardly thickening wedge of barren Tertiary volcanic and sedimentary rocks up to 600 m thick. Brown biotite alteration and copper mineralization are strong, with chalcopyrite/pyrite ratios commonly over 1/1. The zone extends to a depth of at least 1200 m below surface; some drill holes bottomed in strong mineralization. Later, weak to moderate sericitic alteration removed some chalcopyrite and locally remobilized gold. The Pebble deposit has a Measured/Indicated resource of 3 billion tonnes grading 0.28% Cu, 0.32 g/t Au and 0.015% Mo (= 0.5% copper-equivalent) and an Inferred resource of 1.13 billion tonnes grading 0.30 g/t Au, 0.27% Cu and 0.015% Mo (0.55% copper-equivalent). The Pebble East zone has an Inferred resource of 2.13 billion tonnes grading 0.55% Cu, 0.34 g/t Au, and 0.037% Mo (0.97% copper-equivalent). The Pebble East zone is open to the east and to a lesser extent to the north, south, and to depth.

May 1, 2006

By R. Cotton

"1. INTRODUCTIONTo say that the environment will be a dominant issue in the 1990' s is an understatement. A world-wide ""green revolution"" is taking place not only with the general public and politicians, but in the boardrooms and plant facilities of corporations around the world. There is probably not a plant manager who has not been requested by his corporation to assist in some way on environmental issues by providing advice and strategy, by dealing with customer or community concerns, by considering environmental matters in transactions, by developing new environmental compliance policies and requirements, and so on.There are, in the author's view, three main challenges for the mining industry in the environmental area in the 1990's. This paper will discuss each in turn. The first challenge is the concept of ""sustainable development"". Whatever the meaning of this much-debated phrase, we are rapidly moving beyond the buzzword, and into the specifics. It will clearly impact on the way we manage our companies, our communities and our personal lives.The second is the challenge of personal liability; particularly for officers, directors, supervisors and managers of the corporation. For example, in Ontario today, a good percentage of prosecutions for pollution offences are against individuals, usually managers and supervisors, in addition to the prosecutions against the corporation itself for the same offence. In all likelihood, prosecutions will increase in number and will increasingly be directed at individuals. A corporate president has already served a jail term for contempt of court related to his corporation's breach of a court order related to sewer by-law offences. Although the case is on appeal, another president has become the first officer in Canada to be sentenced to jail for violation of an environmental protection statute. Other corporate officers and directors have recently been charged with breach of their duty under Ontario's Environmental Protection Act."

Jan 1, 1991

By K. L. C. Gonçalves, V. K. Alves

"CVRD’s Department of Development and Technology – GETEK has been developing and perfecting a general methodology for scaling up flotation and grinding circuits over several years. This methodology has been applied in several prospective CVRD mining projects. It encompasses a procedure for variability analysis based on bench scale tests, pilot scale grinding and flotation tests and modeling and simulation for scaling up the industrial flowsheet.The bench scale tests include the determination of the Work Index, simplified DWT tests, rougher flotation and locked-cycle tests with samples that are representative of the ore body variability. After completion of these studies, continuous, pilot-scale flotation tests are carried out at a Mini Pilot Plant with drill core samples representative of different stages of the mine life. Finally, studies are carried out with large volume samples using a conventional grinding and flotation pilot plant. This sample is also used for liberation characterization, grinding kinetics, flotation kinetics and industrial circuit scale up. This methodology has been successfully employed in the development of CVRD’s new mining projectINTRODUCTIONCVRD has been developing and perfecting a general methodology for scaling up flotation and grinding circuits over several years. The definition of the processing route and the sizing and scale up of the processing system are some of the products generated during the pre-feasibility of a greenfield project. This paper is focused on the scale up of grinding and flotation circuits, operations known to be of the most importance in a mineral processing plant.The methodology for the definition of the flotation and grinding circuits is divided in the following stages: bench scale ore variability studies, pilot plant operation and modeling followed by simulation. The product of this work is a calibrated simulator that can be utilized for plant operation as well as mine planning, looking to achieve optimal operational conditions.The current models used for scale up of the grinding circuits are aimed at the generating information about the size distribution of the circuit product. However, it is well known that the liberation state of these products may be of critical importance for the performance of the concentration processes downstream. Nevertheless, CVRD has been developing a methodology for liberation kinetics during grinding, aimed at optimizing grinding circuit throughput together with downstream flotation recovery."

Jan 1, 2006

"Mineral separation by exploiting the differences in mineral or metal specific gravities (sg) has been used since antiquity. In the modern era, this is most commonly seen in the recovery of gold by the use of centrifugal concentrators and the separation of coal (low sg) from waste rock (high sg) using heavy media cyclones and jigs. Other metals, notably oxides of tin and tungsten, are successfully recovered using some combination of jigs, spirals and shaking tables. However, many conventional gravity separation systems are limited to fairly coarse (>1 mm) to medium (>100 micron) particle separations due to the limitations of the gravitational separation forces in conventional equipment.Many of the high specific gravity minerals of interest are friable (e.g. cassiterite, wolframite, etc.) and excessive fines are generated in the grinding/comminution circuits which cannot be recovered with conventional equipment and may be lost directly to the tailings stream. Recent developments in centrifugal gravity separation equipment specifically designed for fine or ultra-fine particle separation and recovery have resulted in heavy mineral particle recoveries as fine as 10-20 microns. Ongoing research work has also identified the benefit of closed-circuit, multi-stage gravity recovery circuits using some combination of centrifugal concentrators and conventional equipment (tables, spirals, flotation, etc.) which can substantially increase overall plant recovery and final concentrate grade.This paper will provide details of the evolution of non-fluidized bed gravity concentrators for high sg mineral recoveries, correlate laboratory scale development work with plant results to confirm scale-up, and show how recent flowsheet developments have contributed to significantly improved plant performance.IntroductionGravity concentration is one of the oldest and most cost effective metal and/or mineral recovery systems, ranging from the classic hand panning of alluvial gold, through to the large scale production of a wide range of metals and minerals, including gold, silver, PGMs, black sands, coal and iron oxides. Conventional gravity recovery systems, comprised of some combination of sluices, spirals, jigs and shaking tables have limited application in fine particle (i.e. <100 microns) recovery, or where there are limited differences in particle specific gravities (sg), e.g. the separation of high sg, high value metal oxides (e.g. those of tin, tantalum or tungsten), from medium sg, lower value metal sulfides such as pyrite and pyrrhotite. These unit operations tend to have limited capacity, and while double and triple turn spirals and multi-deck tables have been developed, extensive floor space is often required in order to process the tonnages incorporated in the modern mineral processing plant."

Jan 1, 2017

By Paul Girodo

Government Approval For Foreign Workers 1.Service Canada Work Permit Tasks: i. Prove there are no Canadian Journeymen; ii. Use heavyweight stakeholders; iii. Understand the FWU transformation taking place today. Government Approval For Foreign Workers 2.Provincial Nominee Program (PNP) i. Build relationships; ii. Use heavyweight stakeholders; iii. Continually demonstrate benefits to the Province from foreign Journeymen. Government Approval For Foreign Workers 3.Letter from Union Supporting Temporary Foreign Workers i. Easiest path to approval ?CLAC model

May 1, 2006

By Rimas Pakalnis

This paper describes a focus of work presently being conducted at the Rock Mechanics Research Group at the University of British Columbia. The underhand method under consolidated fill ensures a high recovery under an engineered back that is comprised of cemented rock fill and/or cemented paste fill. This method of mining is generally necessary either due to a weak rock mass comprising the immediate back and/or high induced back stresses. A major concern in the design of sill mats is the loading and strengths associated with the overlying sill mat. This paper reviews past practice coupled with present observations and measurements from over ten(10) mines throughout North America. It outlines areas of concern in terms of design requirements.

Apr 1, 2005

Ontario Operaations Mining Garson Stobie Copper Cliff North Copper Cliff South Creighton Mcreedy East Milling Clarabelle Mill Smelting Smelter Matte Processing Anode Casting Refining Copper Cliff Nickel Refinery Port Colborne Cobalt Refinery

May 1, 2007

By R. Grant, C. Bazin

"Brunswick Mining has investigated grinding media optimization for the purpose of improving the flotation feed and regrind product size distributions to enhance metal recoveries. This paper presents the findings of on-going testing of various sizes and shapes of grinding media. Plant observations combined with estimated breakage rates demonstrate the importance of appropriate selection of media sizes on the milling operation.IntroductionBrunswick Mining and Smelting is located near Bathurst, New Brunswick, and operates one of the world&apos;s largest lead and zinc mines. The ore mineralogy and processing flowsheet have been described elsewhere (Petruk and Schnarr, 1981; Shannon et 01., 1993). Despite the 25 years&apos; milling experience, there has been considerable disagreement as to the required grind size (Grant et 01., 1991). Since 1964, fineness of the flotation feed ranged from 60% to 80% passing 400 mesh (37 um) and averaged slightly above 71% in 1992. Results from mineralogical analysis (Petruk and Schnarr, 1981), laboratory and pilot plant experiments, as well as plant sampling (Grant et al., 1991; Hendriks and Ounpuu, 1985; Shannon et al., 1993) indicate that improved metal recoveries can be achieved by an increase in the fineness of grind, and more specifically by maximizing the amount of material in the 15 urn to 25 um range.This paper shows the results obtained so far from the evaluation of the effect of grinding media size and shape on grinding mill efficiency. The problem of finding optimum grinding media size, shape and composition has been recogrtized since the early 19405. The main criterion for the selection of media size is the size of the material to be ground (Bond, 1958). Although several relationships have been proposed to assist in the selection of grinding media size, none of them are entirely satisfactory (Wills, 1988). Other approaches using the breakage rates and process simulators were developed to help in the selection of grinding media size (Austin et al., 1976; Lo and Herbst, 1986; Concha et 01., 1992; Samskog et al., 1990; Martinovic et al., 1990). Although considerable effort was put into the analysis of grinding media size, very little information is available on the effect of grinding media shape (Norman and Decker, 1985, p. 3C-4). Norris (1952), Kelsall et al. (1973) and Herbst and Lo (1989) have shown that balls are more efficient than other shapes of grinding media including truncated cones and cubes."

Jan 1, 1994

By L. W. Vigrass

Porous and permeable rocks deeply buried in sedimentary basins contain immense volumes of hot water with the potential to supply low-grade heat energy. A part of Canada with sizeable potential for low-grade geothermal energy is the northern Williston Basin of southern Saskatchewan. In this region, the University of Regina has conducted a geological study investigating the feasibility of producing large quantities of warm water suitable for space heating in the Regina - Moose Jaw area. The study has shown that the most favourable geothermal reservoirs are in the Basal Clastic Unit (Cambrian Ordovician), which, at an approximate depth of 2,000 metres has predicted temperatures of 71°C to 74°C. We anticipate sustainable flow rates from these reservoirs of 100 to 300 ms/hr. The water contains about 150,000ppm of total dissolved solids, mostly sodium chloride. It is anticipated that the sulphate content will be 3,000 to 6,000 ppm, that the dissolved oxygen and hydrogen sulphide content will be small, and that the waters will be neutral to slightly alkaline. It is anticipated that the piezometric level will be 46 m below the ground surface for reservoirs in the upper part of the Basal Clastic Unit and 141m below the ground surface for the lower reservoirs of the unit. Plans are to build a low-grade geothermal research, testing and demonstration facility at the University of Regina. To be a practical demonstration, the facility will have to handle very large volumes of water. Because of high salinity, the water will be re injected into a subsurface reservoir. We do not anticipate insurmountable problems because of salinities or as a result of corrosion provided that the waters are not aerated during handling

Jan 1, 1979

By D. Lindeman, R. M. Nendick

"From a technologist's point of view, there have been few fundamental 'breakthroughs' in the process of gold extraction since the turn of the century. However, the industry has made a great deal of progress in its response to the problems that have arisen: a changing demand for products, lower grade and more refractory ores, inflated labour and materials cost and the increasing control of governmental legislation. The trend in North America and South Africa to alleviate these problems has been to adopt a common policy to process larger tonnages of lowgrade ore through more efficient milling circuits. Here, however, is where the similarity ends.The purpose of this paper is to emphasise the different principles of practice employed by extraction plants in North America and South Africa to achieve the same common goal, namely gold bullion.2.0 INTRODUCTIONTwo very important aspects of extraction metallurgy for which the metallurgical engineer is responsible are plant design and plant commissioning.In North America and South Africa the importance of these aspects is due to the fact that both are related to the financial success of the project of which they form part. Good design is imperative if the capital invested in both the plant and the whole mining undertaking is to reap maximum profit. Furthermore, good design leads to a carefully planned and well executed start-up procedure resulting in bringing the plant into production on time, at designed efficiency and within budget.Due to the competent nature of the quartz pebble conglomerate and quartzite ore on the South African Witwatersrand, the characteristics of the ore type are not much different from one mine to the other. Thus a significant operating knowledge and experience has been built up over the years and there is considerably less risk in selecting a gold extraction process route."

Jan 1, 1992

By L. C. St. Arnaud

"Open pit mines can produce large disturbances within regional hydrogeologic systems. The degree of disturbance and the consequences on the local environment depend on the size of the pit and on the physical setting of the mine. Comparisons and contrasts are presented for three Noranda base metal open pit mine sites.The Brunswick Mining and Smelting No 6 open pit has a surface area of 11 ha and a depth of 185 m. It is located within complexly deformed schists and meta-volcanic rocks. The massive sulphide and iron formations visible along the north and southwest walls of the pit generate acid drainage. Mining ceased in 1983, and the pit is presently being filled with waste rock and neutralized drainage water. Water balance calculations estimated a total fill up time of sixteen years. As the water level rises, leakage to groundwater could take place across a fault zone intersecting the pit.The Mattabi Mines F-Group pit has a surface area of approximately 3 ha and a depth of 50 m and is also set in meta-volcanic rocks and acid generating iron formations. Production at this site ceased in 1982, and the low pH pit water is neutralized and pumped into a nearby lake at least every five years. Major water inflows are from precipitation and groundwater. Groundwater mounds within adjacent high land surrounding the pit provide a natural hydraulic barrier to groundwater flow, except for the southeast side. As a consequence, the pit water elevation is being kept below the elevation of the nearby creek in which some of the pit water could drain.The Bell Copper open pit covers an area of 40 ha and is 270 m deep. With present ore reserves, the mine might be in operation until 1992. The geology around the pit consists of andesitic tufts, siltstones and shales, and the acid generation potential of the rock is relatively low. A recent hydrogeological study estimated fill up time by precipitation and groundwater to be in the order of 260 years.With increasingly more stringent requirements for mine closure, it is anticipated that more detailed hydrological and hydrogeological studies of open pits will be required to determine post-closure environmental impacts and the suitability of open pits for mine waste disposal."

Jan 1, 1990

By N. W. Johnson

"Two analysis tools which have been applied to problems in industrial flotation processes are the application of kinetics and detailed mineralogical or liberation data, both on a size-by-size basis. This paper describes practical examples of the application of kineticsto examination of the entrainment recovery mechanism, the retardation of valuable minerals at the commencement of flotation sections and the resolution of problems related to valuable species with reduced rate constants at the end of sections. Examples of the usefulness of size-by-size liberation data for monthly composite samples of products from a concentrator are also described.IntroductionMuch of the published industrial flotation data on the behaviour of minerals in various size fractions have been obtained from specially conducted surveys with a duration of several hours. Such data are very valuable and provide a record of the behaviour of the entire flotation plant or sections during the interval of the survey. Some examples from the literature can be cited Cameron et al. (1971), Lynch et al. (1981), Frew (1982). Less attention has been given to the value of complementary sized data derived from compositing shift or daily inventory samples over an interval of one or several weeks with repetition of the procedure at the completion of each interval. Mineralogical point count data of the size fractions of such samples and data interpretation are rare in the literature.Kinetic data on the behaviour of valuable and gangue minerals in various size fractions have been reported often for flotation banks. In the experience 0 f the writer, there has been a tendency among operating metallurgists to ascribe problems in the flotation process to rate constant values for gangue minerals which were believed to be unsatisfactorily high. Less attention has been paid to the value for the rate constant for the valuable mineral over the length of a particular bank.Examples of the types of data which were mentioned previously have been obtained from the lead-zinc concentrator at Mount Isa Mines Limited. The various types of data have been collected to evaluate their worth in order to provide relevant information for both research and operational aspects of the plant."

Jan 1, 1987

By C. A. Olivier

Le Ministère des Richesses naturelles du Québec has continued development work on the process/or direct reaction 0/sodium carbonate with lithium-bearing silicates(Canadian Patent No. 643,843). Recent process improvements have involved continuous pressure leaching, rather than batch leaching, and higher solution concentrations (over 55 g/l Li2CO3. These improvements have been demonstrated on a pilot scale operating on decrepitated spodumene from the Sullivan Mining Group&apos;s Quebec deposit. The new process is expected to be very competitive with the conventional sulphuric acid route. The improved flowsheet and main process steps will be described along with pilot-plant results.

Jan 1, 1979

By E. Lampard

Mines must manage variability from within their operations as well as those imposed upon them by external forces. This paper provides a framework to manage both internal and external variability through the review and adjustment of the mine plan where applicable to minimize lost production and maintain operational safety. KEYWORDS Mine plan, mine schedule, variability, safety, lost production, float days, update, adjustment, operations, Maintenance

May 1, 2011

By A. Sheikh, S. Hardcastle

"Various methods of leaching are being used by Denison Mines Ltd. Uranium Operations. This study was designed to investigate the flood leaching operation of two parallel inter-connected stopes contained by pressure bulkheads. The teaching muck piles, created from large scale uppers production blasts, are aerated by compressed air lines laid beneath the pile prior to blasting.A SFr, tracer gas, injection technique has been employed to determine the post-blast integrity of the single-ended air distribution system. SF6 was injected into the compressed air line feed to the stopes and collected at borehole exhausts from the stopes in grab-sampling bags. Analyses of the air samples were performed with a gas chromatograph.The system was designed to give a sufficiency of oxygen to sustain bacterial activities used to enhance the leaching recovery in the muck piles. The results obtained indicated that there is a minor anomaly in the distribution system that may not be critical to the operation. Compared to an idealized case, the over-all efficiency of aeration throughout the muck is estimated at over 85%."

Jan 1, 1988

By E. T. C. Spooner, Ioannou S. E.

"Abstract- The Dixie Zone Au-Ag deposit is a low sulfidation (adularia-sericite) epithermal system located 1.4 km from the producing Ken Snyder mine (7.0 M oz Au equivalent reserve/resource), Midas Mining District, north-central Nevada. This WNW striking, steep south dipping, fault system is hosted by mid-Miocene (17 Ma to 14 Ma) felsic volcanic rocks that were erupted from a vent(s) to the northeast. The system was mineralized by approximately 250 ± 20°C (n = 86), near-neutral pH, ascending fluids of predominantly meteoric origin [volatile corrected salinity = 0.52 ± 0.35 eq.wt% NaCl (n = 86)]. Bulk analytical fluid inclusion gas chromatography identified a mineralizing fluid volatile composition of 99.68 ± 0.18 mole % H2O, 0.16 ± 0.09 mole % CO2, 0.11 ± 0.05 mole percent N2, 0.05 ± 0.04 mole % CH4, and 0.01 ± 0.01 mole % CnHn+i (n = 7). Sulfur species were not analyzed due to absorption within the chromatography system; however, a H2S concentration of 16 ppm to 160 ppm molar is estimated. The volatile content contributes approximately 50% of the microthermometrically measured freezing point depression. Cl- concentrations up to approximately 9000 ppm suggest an additional, probably magmatic, fluid component.Ascending fluids were channelled along approximately NW striking faults related to the same primary extension (17 Ma to 14 Ma) responsible for Nevada basin and range physiography. Boiling induced by CO2 effervescence led to precipitation of Au and Ag over a 150 m vertical interval. At the time, this interval was about 460 m below the water table (approximately 38 bars), but because of erosion, it is now within 90 m of the surface. A Ag:Au ratio of 25:1 has been calculated for the Dixie Zone."

Jan 1, 2000

By D Reddy Chinnasane

Outline Location of BirchtreeMine and Layout of 124 OB Cut and Fill Mining Background Purpose of the Studies Methodology Results and Analysis Conclusions and Recommendations

Oct 1, 2010