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By PáDeáBosscher, A Bruwier, M Weidenbach, C Philippe, T Ly, A Van Den Bosch

In the mining business the cost of grinding media represents up to 40 per cent of the total milling cost and therefore grinding media optimisation has a direct impact on the concentratorsÆ bottom line. In the process of optimising grinding media alloys, Magotteaux developed its own methodology which consists of a series of analyses, tests, calculations and evaluations and assist the concentrator in optimising their processes.This paper presents the Prominent Hill case study where this methodology has been applied on both the secondary ball mill and regrind IsaMill«.The analysis of the wear conditions in the secondary ball mill allows Magotteaux to select a series of grinding media alloys to be tested to reduce grinding media consumption. By marking these different grinding media types with a unique drilling pattern, one can quantify the relative wear performance and cost effectiveness of each grinding media type within the same milling environment. After selecting the most economical grinding media composition a full scale industrial trial, involving topping up the selected grinding media is carried out and the media wear performance closely monitored and analysed.For stirred grinding mills, suppliers propose different grinding media, mainly ceramic beads, with specific properties and cost. Therefore the concentratorsÆ choice is a compromise between grinding efficiency, wear resistance and cost. The selection of the correct grinding media for stirred milling and the assessment procedure will be presented in this paper. The test results of the new stirred grinding media type in the regrind IsaMill« at Prominent Hill are presented and analysed.CITATION:Weidenbach, M, Van Den Bosch, A, Bruwier, A, Ly, T, De Bosscher, P and Philippe, C, 2012. Forging ahead - A collaboration between the OZ Minerals Prominent Hill Concentrator and Magotteaux to improve plant performance, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 105-110 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

By Boris Chacón

Durante el desarrollo cotidiano de una operación minera, se obtienen diversos tipos de información como: muestreo, perforación diamantina, etc. las cuales, se originan en base a observaciones de campo. En ese sentido, es vital desarrollar procedimientos adecuados que permitan administrar la información compilada. Por ello, el presente trabajo señala la aplicación de la administración de datos geológicos en la empresa minera Iscaycruz. La información de datos geológicos se puede administrar en formatos tipo gráfico o en bases de datos, por lo cual se requiere que la estructura de almacenamiento de dichos formatos contenga parámetros previamente analizados. Asimismo, es necesario que el equipo de trabajo organize la base de datos, a través de los objetivos y metas trazadas, así como de la verificación y análisis de los resultados, para mantener actualizada la información.

Aug 25, 2002

By G J. Lyman, F S. Bourgeois

As the prices of metals slide, financial control of mineral processing operations becomes more and more important. Control of a mineral processing operation implies measurement of grades, recoveries and losses, with every effort being made to recover more product from the same feed. Shareholders become more concerned with the transparency of the operations and the corporate governance.There is only one way to achieve this and this is by sampling and making good use of the results of that sampling. Sampling systems that provide only 'rough' figures are not good enough. Similarly, a competitive position in the marketplace demands close control of product quality. Maximisation of product output demands that the product just meet the specification so that product yield is maximised. Decisions based on precise sampling results carry less risk that those based on imprecise results.This paper reports some recent advances in sampling theory that permit full quantification of sampling distributions, not just the sampling variance as has been possible in the past. These advances can be used together with simulation methods to tie down sampling precisions and develop calculations that quantify decision making risk when using the sampling results. A number of examples illustrate the use of the new advances and a final example indicates the development of risk analysis curves in the coupling of sampling precision to sale of a commodity.CITATION:Lyman, G J and Bourgeois, F S, 2015. Sampling, corporate governance and risk analysis, in Proceedings MetPlant 2015, pp 27–34 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 7, 2015

By S. M. Moore

Each year more than 400 coal miners are injured (fatal and non-fatal) by rock falling from between or around roof supports. Many of these injuries could be prevented by the installation of roof screen (wire mesh). However, many coal mines may be reluctant to use roof screen because of the added cost. The goal of this study was to determine the potential saving in workers? compensation (WC) premiums that could be achieved due to a reduction in rock fall injuries after roof screening. The WC rate-setting methods utilized by IL and KY were investigated in this study. Using actual data obtained from MSHA, national and state WC bodies and individual insurance companies (e.g. average cost per injury, loss cost rate, number of injuries per year, number of injuries preventable each year with roof screening), hypothetical mines (representing two mine sizes: 67 and 150 employees) were constructed with realistic ranges for estimates of injuries and WC premium costs. Using each state?s actual WC rate-setting formulas, total savings in WC costs after a three-year period were determined. Across both states, savings in WC premiums ranged from 1.8% to 14.6% when injuries were reduced by 10% to 20%. Additionally, an economic analysis was performed for an existing mine consisting of 67 employees. For this mine, the annual cost of a roof screening program was estimated at $240,000. If the screening prevented 13 ?struck by? WC claims over a 3-year period at this mine, the reduction in WC premiums alone could pay for the entire screening program.

Jan 1, 2010

By N. Dashbold, E. Rudnik

"Waste electric and electronic equipment currently comprises the fastest-growing waste stream in the world, but at the same time it is seen as an important source of metals for the recycling industry. E-waste is usually treated pyrometallurgically, with hydrometallurgical methods used to a lesser extent. This paper reports the results of research on the selective recovery of copper from smelted low-value electronic waste. Pyrometallurgical pretreatment of the scrap allowed the removal of plastics and the increase of metal content in the material. The obtained alloy of copper, zinc, tin and silver was a multiphase solid consisting of two brass phases and inclusions rich in iron, lead and silver. Copper alloy was further anodically dissolved in ammoniacal chloride solution. It resulted in high degradation of the material and accumulation of the metals mainly in the slime. The slime was then leached in acid or ammoniacal chloride and sulfate solutions followed by selective copper electrowinning. Hydrochloric acid was the most efficient solvent for the slime, but ammoniacal solutions were more selective for copper. Copper could be leached with 96 to 100 percent and 87 percent efficiency from the slime by the chloride and sulfate solutions, respectively. Copper with 90 to 99 percent purity at current efficiency of 42 to 76 percent was obtained from the acid solutions, while copper with 98 to 99 percent purity at current efficiency of 60 to 86 percent was deposited from the ammoniacal baths. IntroductionObsolete electrical and electronic equipment makes up the fastest-growing waste stream in the world (Baldé et al., 2015), due not only to the introduction of wider ranges of new types and models of a variety of electronic products but also to the shortening of their life-spans, sometimes referred to as planned obsolescence by producers (Keeble, 2013). About 42 million tons of electronic waste, or e-waste, are estimated to have been generated globally in 2014 (Baldé et al., 2015), and it is expected that in the European Union alone the waste electrical and electronic equipment (WEEE) stream will reach more than 12 million tons by 2020 (European Commission, 2015). Accurate evaluation of e-waste generation (Wang et al., 2013) and its global transport (Breivik et al., 2014) is difficult, mainly due to a lack of high-quality data related to its market and socio-economic dynamics.Many countries introduced legislation to control e-waste flow, collection and recycling. For example, Directive 2012/19/EU on waste electrical and electronic equipment (European Parliament and the Council of the European Union, 2012) obliges all E.U. members from 2016 to maintain a minimum collection rate of 45 percent of the total weight of WEEE collected. From 2019, the minimum collection rate to be achieved annually shall be 65 percent of the average weight of electronics placed on the domestic market in the three preceding years, or alternatively 85 percent of WEEE generated in the territory of the particular member."

Jan 1, 2017

By Sun Jiale

This paper presents a brief description of information construction technology for supported excavation to assure satisfactory performance of engineered system. As a successful example of application of the technique, quite a few troublesome problems, such as high groundwater level in soft ground and crowded buildings around the excavation, were solved satisfactorily in the construction of the new tower of Guangzhou Overseas Chinese Manison.

Jan 1, 1991

By Jan Kvinge, Aina Opsal Bakke, Yanjun Li, Jan-Ove Løland, Svein Jørgensen

"Compound casting is a very attractive approach to reduce the weight of components for some specific applications while keep both the functional and the mechanical property requirements. In this work, Al-7Si-Mg alloy /copper compound castings were produced through a low pressure die casting process. To reduce the oxide present in the interface, flux coating of the copper pipes was tested. The interface layer formed between the two alloys was investigated using Optical Microscopy (OM), Scanning Electron Microscopy (SEM) and Energy Dispersive X-ray Spectroscopy (EDS). Vickers Micro-hardness was also measured across the interface. Results showed that the flux coating prevented the formation of the metallic bond between copper and aluminum. Instead, high concentrations of potassium, magnesium and fluoride, coinciding with formation of KMgF3 and MgF2, were detected in the interface. In the castings without surface treatment, a continuous metallurgical bond was achieved. In the bond layer, various Al-Cu intermetallic phases have been detected. In addition, formation of primary silicon particles in the reaction layer was observed in all castings. The mechanism for the formation of the intermetallic phases and the strength of the interface layer have been discussed.INTRODUCTION Recent development in the automotive industry has focused on producing lightweight components, which can reduce CO2 emission and increase the efficiency. Often, one material alone is not able to meet the requirements for a specific use. The need for lightweight bimetallic components has therefore increased. Aluminum-copper bimetallic components can be used in wires and bus bars where conductivity is an important factor (Lee, Bang, & Jung, 2005). Compared to Al-Cu alloys, Al-Cu bimetallic components can reduce both weight and cost without reducing electrical and thermal conductivity (Sheng, Yang, Xi, Lai, & Ye, 2011). A challenge is, however, that aluminum and copper have high affinity to each other, especially at elevated temperatures. This causes formation of brittle intermetallic phases with high electric resistance (Liu, Wang, Sui, Wang, & Wenjiang, 2016). In addition, both metals are subjected to oxide formation on the surface. Copper is especially exposed to oxidation at elevated temperatures (Pinnel, Tompkins, & Heath, 1979), while a thermodynamically stable oxide layer will spontaneously form on the aluminum surface (Papis, Loeffler, & Uggowitzer, 2009). Oxides are known to be passive substrates, which will reduce wettability and will thus prevent the formation of metallic bonding between aluminum and copper (Papis, Hallstedt, Löffler, & Uggowitzer, 2008)."

Jan 1, 2018

Situated in the Greymouth Coalfield, the Mt Davy Block contains a significant resource of high quality bituminous coal suitable for use in metallurgical coking coal blends. The coal, which is all low sulphur, is seen as a complementary deposit to export coking coal at Stockton Mine.   Previous drilling by State Coal Mines (1948 - 1953) and Japanese interests (1970's) have been supplemented by Coal New Zealand drilling (1992 - 1993) and surface mapping. A mining feasibility study was completed in late 1993 and approval gained for the driveage of an exploratory tunnel to the target seam.   Following construction of surface facilities, tunnel driveage on a 1:5 decline commenced from Rewanui in May 1995 and to date about half the required length has been driven. When the seam is intersected a comprehensive geotechnical testing programme will be undertaken to assess the in-situ mining conditions. It is expected that the results will confirm the assumptions made in the feasibility study.   The Mt Davy Mine, at depths of 500 - 700 metres, will be the deepest underground coal mine ever inNew Zealand and in-situ stresses (orientation and magnitude) will be important considerations for mining engineering. The coal has a high gas content and suitable techniques for gas control and prevention/mitigation of coal outbursts will be necessary.   Mine development is scheduled to commence in mid 1998 and an annual production level of 500 000 tonnes is forecast. The coal seams will be developed by continuous miner and coal extraction will be by hydraulic monitor. Coal will be pumped to the surface and from Rewanui transported five kilometres to Dunollie either by truck or slurry pipeline. At Dunollie a coal dewatering plant, coal stockpile and rail loadout will be constructed.

Jan 1, 1996

By Fred Breu, Ronald E. Heuer

INTRODUCTION The Point Lepreau nuclear generating station was constructed in the late 1970's and early 1980's by the New Brunswick Electric Power Commission. Point Lepreau is a small peninsula on the north shore of the Bay of Fundy, about 55 km west of St. John, New Brunswick. Cooling water for the plant was to be obtained from and discharged into the Bay of Fundy by the system of shafts and tunnels illustrated in Fig. 1. The design engineer for the cooling water system was Albery, Pullerits, Dickson and Associates, Ltd., Toronto. Ferrera-Resco Ltd, Windsor, Ontario, was contractor for construction of the cooling water shaft and tunnel system, which was done between 1976 and 1979. The senior author was project manager for Ferrera-Resco, the junior author served as consultant to Ferrera-Resco on evaluation of certain geotechnical questions which arose during construction. SITE CONDITIONS Geology Except for limited overburden soils at the onshore shaft sites, the tunnels and shafts were constructed entirely in Triassic conglomerate, sandstone, and siltstone, with some shale interbeds. The bedding strike averaged about N6°E, The dip was to the west at about 30° along the Discharge Tunnel and 40° along the Intake Tunnel. Rock along the Intake Tunnel was mostly sandstone and siltstone and more thinly bedded, with RQ D generally in the range of 50 to 60%. Rock along the Discharge Tunnel tended to be coarser sandstone and conglomerate and more thickly bedded, with higher RQ D averaging about 80%. In addition to bedding joints, two other joint systems were identified

Jan 1, 1985

By P Dowd, R Jeuken, C Xu

In Australian coalmines there are a number of coal quality parameters that are measured by analysing samples from boreholes. The data points are used to generate spatial models of these parameters, typically using inverse distance as an interpolation method. Numerous additional boreholes are drilled which have geophysical logs measured downhole but from which no coal quality samples are taken. Full suites of coal quality analyses are expensive and time consuming to drill and sample. If geophysical proxies for coal quality could be used, this would greatly improve the accuracy of the resource estimation without additional cost.This paper examines two of the most important coal quality parameters – ash and volatile matter – and assesses the results of a number of geostatistical interpolation methods. The aim is to improve the accuracy of the interpolated models. Auxiliary variables from geophysical logs are also trialled to improve the estimation.A multivariate analysis of ash, volatile matter and several auxiliary variables of a selected coal seam was undertaken in a spatial context, using variography. This was used to construct a model of co-regionalisation that relates all of the variables as manifestations of an underlying geological theme.The traditional inverse distance interpolation approach was compared to ordinary kriging, universal kriging, co-kriging, regression kriging and kriging with external drift approaches. The relative merits of the six interpolation methods were compared by removing all points, one at a time and estimating at the thus known location. The mean error and the root mean square error were evaluated as a measure of bias and precision.This paper demonstrates that there is significant opportunity to improve the estimations of coal quality when using kriging with external drift. For example, the results show that when using the depth of a sample as an external drift variable there is a significant improvement in the accuracy of estimation for volatile matter. Furthermore, when using borehole wireline density logs as an auxiliary variable for ash content, there is improvement in the estimation of the in situ ash. The underpinning reasons for this are discussed in the context of the underlying geology and the physical properties of the coal. These findings hold promise for utilising cheaper proxies for coal quality parameters to dramatically improve data density and the quality of estimations.CITATION:Jeuken, R, Xu, C and Dowd, P, 2017. Improving coal quality estimations using geostatistics and auxiliary variables, in Proceedings Tenth International Mining Geology Conference 2017, pp 157–168 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 20, 2017

"INTRODUCTION Early in 1936 the Interstate Oil Compact Commission and other groups interested in forecasts of demand for petroleum requested the Bureau of Mines to make a physical inventory of crude-oil stocks in the United States, Accordingly, the 74th Congress appropriated funds for conducting such a survey during the fiscal year ended June 30, 1937. In the discussions and hearing 2/ relative to the need for the survey, the opinion was expressed that crude oil held in storage for long periods had lost so much of its gasoline content that the economic significance of total stocks should be discounted materially. Therefore, a primary objective of the survey was to make a physical inventory of the crude oil held in storage above ground and to determine the quantities and characteristics of the light fractions contained in the various stocks."

Sep 1, 1938

By Gregory M. Molinda

The Coal Mine Roof Rating (CMRR) was developed 10 years ago to fill the gap between geologic characterization and engineering design. It combines many years of geologic studies in underground coal mines with world-wide experience with rock mass classification systems. Like other classification systems, the CMRR begins with the premise that the structural competence of mine roof rock is determined mainly by the discontinuities that weaken the rock fabric. However, the CMRR is specifically designed for bedded coal measure rock. Since its introduction, the CMRR has been incorporated into many aspects of mine planning, including longwall pillar design, roof support selection, feasibility studies, extended-cut evaluation, and others. It has also become truly inter-national, with involvement in mine designs and funded research projects in South Africa, Canada, and Australia. This paper discusses the sources used in developing the CMRR, describes the CMRR data collection and calculation procedures, and briefly presents a number of practical mining applications in which the CMRR has played a prominent role.

Jan 5, 2007

By T. Fabritius, L. Hekkala, J. Härkki

The Outokumpu steel belt sintering technology is used for manufacturing chromite pellets that are charged into smelting furnace for ferrochromium production. Hot air is blown through the moving bed of raw pellets. Direct temperature and flow measurements are difficult to do in the bed of a real scale belt sintering process. Therefore, a tested and reliable mathematical model would be a valuable help for optimising operational parameters of the process and designing new constructional improvements. The aim of this work is to present a mathematical model that calculates the gas flow and temperature distribution in the bed of pellets and the atmosphere. Computational fluid dynamics program FLUENT is used to calculate the gas flow, composition and temperature of the gas in the sintering bed during processing. Energy equation for the bed, containing the effects of evaporation of water and oxidation of carbon and iron oxide, will be added to the model to obtain the temperature of the pellets. The porous media model used for pressure drop calculations in the bed is shortly discussed. A comparison between CFD-calculations and experiments with a pilot scale batch-sintering reactor are also presented. The final model can be used as an engineering tool when developing the process and the equipment further. Different designs can be easily tested with the model and the effects of the changes to the flow pattern and temperature distribution can be studied.

Jan 1, 2004

By F. Crundwell, N. Steenekamp

"The leaching of sulphide minerals generates a significant amount of energy. There are three systems that might be used to remove this energy. These are: cooling coils, quench water, and flash recycle. Flash recycle is the recycling of slurry from the autoclave to the feed tank. Heat is removed in this arrangement when the slurry at elevated temperature and pressure is flashed to atmosphere, releasing steam. The slurry is returned to the autoclave at lower temperature, resulting in cooling. The cooling is directly proportional to the steam released from the recycled slurry. Design equations and results for the flash recycle system are presented. A fourth proposal for removing heat is to recycle slurry from the back of the autoclave. These four alternatives not only remove the heat, but also have an effect on the leaching efficiency. Amongst the four alternatives, design calculations indicate that the flash recycle option is the optimal solution in terms of leaching efficiency and heat removal, and recycling product from the back of the autoclave is the worst. In spite of its efficiency in terms of heat removal and leaching, the implementation of the flash recycle system is limited to a few operations in the platinum industry.INTRODUCTIONIn considering ways to improve leaching operations, focus is placed on the chemistry, on the kinetics, the agitation and perhaps even on the gas mass transfer conditions. Attention is less frequently placed on the removal of energy from these reactors and how this might limit performance.The pressure leaching of sulphides releases a significant amount of heat. For example, the reaction for the leaching of chalcocite, Cu2S, given by"

Jan 1, 2012

Since the last occasion on which I spoke to you on this subject there has been a considerable increase in the mechanical ventilation of Western Australian mines.The following table compares the figures for 1939 with those of 1941. The type of fan most commonly used is the limit load fan. On account of its higher mechanical efficiency it has very largely superseded the multivane fan with forward curved blades.Two fans of the propeller type are in use and these have given very satisfactory service. This type of fan also has a high mechanical efficiency particularly at the correct duty for maximum efficiency.The maximum efficiencies obtainable from the various types are, for "multivane" 670/0, for "limit load" 75%, and' for aerovane 7810/0.References to propeller fans with variable pitch have appeared in the technical press, but no detailed d~scription has been issued. The advantages of these fans for mine ventilation are not obvious. An aeroplane is required to operate over a wide range of duties, while the speed of the motor for maximum efficiency is pl'actically constant. In such circumstances the propeller can be adapted to the particular work it has to do by varying the pitch of the blades. In a mine the speed of the fan can be varied to suit.the conditions, and it would appear that variable pitch is not necessary while there is a reasonable range of fans to choose from.The question of the correct position for the fan still comes in for considerable discussion. Exhausting fans at the surface have the support of time and custom, and in many cases this represents the only reasonable arrangement. Such fans should be fitted with an evase, especially in the case of fans with a high outlet velocity such as the aerovane types.It is claimed that there are advantages in having the fan at the surface where it is readily available for inspection and maintenance. Experience has shown that these advantages are more apparent than real. Fans will run for long periods with no attention other than occasional greasing and cleaning, and modern electric motors will run without attention for very long periods.In some instances the expense of a power cable to take current to the fan may be saved by placing it on the surface, but in the majority of cases...

Jan 1, 1941

By Arthur E. Wells

"The present and possible future commercial production of potash from various sources in the United Status is discussed below under the following heads:1. Potash from the saline lakes.(a) Nebraska lakes.(b) California lakes.(c) Other lakes.2. Potash as by-product from the cement kilns.3. Potash as by-product from the iron blast furnace.4. Potash from silicate rocks.(a) Feldspathic material.(b) Leucite rocks of Wyoming..(c) Sericite slates of Georgia.(d) Glauconite (greensand) of New Jersey.5. Potash from Alunite.6. Potash from organic sources other than kelp.7. Potash from kelp.In this discussion very little more than passing reference, if any, will be given to those projects which up to date have been developed only slightly, if at all, beyond the small scale laboratory stage, or for the development of which there appears to be very little financial and technical backing at the present time. Nether or not many of these proposed projects have merit can be determined only by an actual operating demonstration.Generally speaking, the potash industry of the United States has been hardly more than marking time since the armistice. This is, largely due to the fact that the consumers of potash have been expecting that heavy shipments of foreign potash to the United States would be resumed soon at somewhere near the pre-war prices, and they have not seen owing domestic potash at present prices for many months. The result has been that domestic producers have been unable to sell their product at even cost and many producers of potash have been obliged to close down. Thu Nebraska lake plants were closed down in February. The California saline plants were operating at below capacity, and the only large alunite plant is closed down. Even the cement plants have little or no inducements to recover the potash being lost in their waste gases and the production is very small at present.This condition will doubtless continue until a definite policy is decided upon as to the future importation of potash salts from Germany and. France. If unrestricted importation is allowed at prices even 50% greater than in 1913, only a very few, if any, of the domestic projects will be in position to compete with the foreign potash immediately, and the present industry will probably greatly decline with little or no prospect for a future resumption of the industry."

Aug 1, 1919

By J. E. House, C. J. Lewis

In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-pregnant sulfuric acid leach solution is contacted with an organic extractant dispersed in kerosene; the uranium enters the organic phase and associates itself with the organic extractants. Simultaneously an-ions or cations, as the case may be, leave the organic extractant phase and enter the aqueous phase. The two phases are then allowed to separate by virtue of their immiscibility. Based on the data available, a generalized procedure for the recovery of molybdenum from the sodium carbonate scrub liquor of acid leach uranium mills using secondary and tertiary amine extractants is presented, as is another generalized procedure for the recovery of molybdenum from the bleed stream of alkaline leach uranium mill circuits. Such economic appraisal as can be made at this stage of the development is also indicated. Molybdenum occurs in most uranium ores. Whether the uranium is solubilized by the acid leach route or by the alkaline leach process, some of the molybdenum present in the ore is solubilized along with the uranium. Molybdenum in uranium mill pregant solutions may cause processing difficulties, among which may be mentioned the poisoning of resins, the precipitation of organo-molybdenum complexes, or the build-up of molybdenum in recycle streams. While the uranium milling industry has been successful in applying measures to overcome the problems arising from the presence of molybdenum in the circuits, such measures nevertheless add to the cost of producing uranium yellow cake concentrate. Table I shows the analysis of typical uranium pregnant leach solutions. Uranium recovery by solvent extraction is presently practiced only in conjunction with acid leach circuits, and it may be noted that of about 16,000 tons of uranium ore daily processed by sulfuric acid leach, more than half of this tonnage follows the solvent extraction route. Uranium mills leaching ore with sodium carbonate-sodium bicarbonate solution (the so-called alkaline leach circuit) do not use solvent extraction. In this case the uranium pregnant solution is sufficiently pure to permit the direct precipitation of uranium concentrate with subsequent recycle of the alkaline leach solution. While solvent extraction procedures may be developed which are economically attractive for the alkaline leach uranium process, this has so far not been the case. The earlier uranium mill solvent extraction circuits used organo-phosphate extraction reagents, among which di-2-ethylhexyl phosphoric acid was popular. It should be borne in mind that uranium in its sulfuric acid leach solution exists in an equilibrium involving uranium in both anion and cation forms: uo + 2S0 = [uo2(so4)2] For this reason uranium may be extracted by either cation extractors or anion extractors and this is likewise true of molybdenum. However, the organo-phosphate extractants also load ferric iron which is usually present in the uranium acid leach solution. For this reason, it is necessary to reduce this solution substantially to the ferrous state ahead of solvent extraction using organo-phosphate reagents. The amines, being anion extractors, selectively reject iron and thus render the iron reduction step unnecessary. In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-

Jan 1, 1961

By William E. Young

There are many large rhodonite-bearing veins containing manganese in the Eureka-Animas Forks area of the San Juan Mountains in Colorado. The Sunnyside vein is the largest known vein in the area. As observed on the surface, the vein is mineralized over 8,000 feet and attains a maximum width of 200 feet. The average width of the vein and its branches and ramifications is 58 feet in the area that was mapped in detail. The weighted-average analysis of surface and underground samples from the northeastern extension of the Sunnyside vein is 8 percent manganese. Most other known veins in the area have a similar manganese content. The Hidden Treasure vein lies southeast of the Sunnyside-No Name-Mastadon vein. The weighted average of the analyses of surface and underground samples from this vein is 13 percent manganese. Exposures of the Scotia-Silver Chord vein are visible over a strike length of 4,000 feet on Treasure Mountain. Diamond-drill holes show that the vein contains 10 percent or more manganese to a depth of 400 feet. The vein has an average width of 28 feet. The Independence vein is exposed on the surface for a distance of 600 feet and has an average width of 6 feet. Three cut samples taken across the vein had an average content of 20.7 percent manganese. Sampling of the outcrops of eight veins in the Placer Gulch area also indicates manganese. Numerous other veins cropping out in the surrounding area are similar to the outcrop of the Sunnyside vein system mineralogically and are worth investigating.

Jan 1, 1966

By John E. Young

U.S. Steel Corp.'s Winifrede mine is operated by Lynch District in the mountainous region of eastern Kentucky. This region is characterized by narrow finger ridges and precipitous slopes, which present both an opportunity and challenge in the design and selection of coal-handling facilities for the movement of coal from the section belts to the railrod cars. Lynch District selected diesel-powered locomo¬tives and 100-ton articulated cars capable of negotiating sharp curves. The drop¬bottom mine cars are equipped with longitudinally hinged doors that are auto¬matically opened and closed. Lynch District selected a large-diameter steel-lined borehole for lowering the coal approximately 1000 ft from the seam level to the valley floor railroad load-out point. Taking advantage of gravity, the presized coal is "lowered" vertically approximately 720 ft to existing entries of a previously abandoned mine. The coal is then conveyed horizontally into storage. The coal¬handling facility from the mine-car dump to the railroad load-out point is fully automatic through use of tone control. TV cameras are installed along the facility to continuously monitor the strategic points. The diesel locomotives are either directly operated or remotely radio-controlled, depending on the functions Per¬formed. U.S. Steel Corp.'s Winifrede mine is operated by Lynch District in the mountainous region of eastern Kentucky. This region is characterized by narrow mountain ridges and precipitous slopes which present both an opportunity and challenge in the design and selection of coal-handling facilities for the movement of coal from the section belts to the railroad cars. The narrow finger ridges precluded use of under¬ground haulage and, thereby, presented the opportunity to use high-capacity equipment on the outside bench. The precipitous mountain slopes, steepest at the Wini¬frede seam horizon, presented a challenge to select a solution for lowering the coal approximately 1000 ft from the seam level to the valley floor railroad load-out point. It was evident from the start of planning that these two factors predominated; namely, rail haulage and coal lowering, and solutions to these would determine the final design of the Winifrede mine coal-handling facilities. The Winifrede mine derives its name from the Wini¬frede Seam which comprises the total reserve recov¬erable from this facility. The seam is 5 ft high with a good siltstone bottom and competent shale or sandstone roof. The seam is at an elevation of 2700 ft where the mountain ridges are long and narrow. This configuration of the reserve was ideal for conveyor belt installation to transport the coal from the working face to the outside bench. The preclusion of underground rail haul¬age permitted an opportunity to review the type of locomotive power and to select high-capacity equipment without the limitations associated with underground systems. Since the basic electrical design of the mine was al¬ternating current, with rectification to direct current for shuttle cars required only in the face areas, there was no point in bearing the burden of additional cost of dc trolley installation. And, since the haulage would re¬quire only a few tunnels to negotiate sharp points and to reach the back side of the reserves, diesel power was selected for the haulage locomotives. To provide stability on the almost continuously curv¬ing track and to provide the maximum width of haul¬age equipment, both locomotives and mine cars, stan¬dard railroad gage was chosen to offer the largest choice of equipment. The decision was made to select a standard 45-ton railroad diesel locomotive but to lower the cab height to reduce tunnel excavation requirements. The mine car was the next challenge. Past experience in Lynch operations indicated that drop-bottom cars would be a successful application with the free-flowing coal. The search was for a drop-bottom car within the dimensions of the locomotive to discharge the coal in the shortest time. Numerous types were considered and visits made to various sites to observe cars in operation.

Jan 1, 1972

By L. H. McKay

Tailing Dams Three types of dam have been structed to impound tailings from the Allenby mill: 1.- By classifier unit 2.- By earth dam 3.- By direct deposition of tailings The first type, which includes two dams, was built across a shallow valley which sloped down to Similkameen river, just below the mill site. Two earth embankments, one below the other, were built across the valley. Tailings were then piped to Akins classifiers mounted on wooden wheels. The classifiers were set up at one side of the valley and gradually moved ahead on the sand embankment created by the sand discharge from these classifiers, the slime overflow going to settli.ng ponds formed by the dams. Decanting wells were built in the settling ponds by placing four six-by-six timbers vertically in the ground, two feet apart, to which two-by-six tongue-and-groove boards were nailed. The elevations of the ponds were raised, as required, by nailing additional two-by-six boards on the well and bolting on more six-by-six timbers when required. The bottoms of the wells were attached to a covered launder made of three-inch lumber laid on a one per cent grade. This launder carried the overflow through the dam. The second type of dam was built across a long, narrow valley, known as Hitchen's Flat, the upper ~n d of which was about 2½ miles north of the mill. This valley, an ideal tailings site, was 4,000 feet long and from 600 to 800 feet wide. An earth embankment with a concrete spillway through the top was built approximately 28 feet high across the lower end of this valley. Tailings were then dumped into the upper end of valley, with no further expense until the basin was filled up to the level of the flume where this entered the valley. When the upper end of the valley was filled to this point there was still room for more tailings at the lower end. Therefore, a pump was installed to elevate the tailings to build a wall around the upper end of the valley by depositing sands from a launder built up on 20 foot supports and sloping both ways from the pumping station around the upper end of the basin.

Jan 1, 1950