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By V. E. Edlund

Laboratory batch flotation tests were conducted on copper converter slags to evaluate the relative merits of recovering copper from slow-cooled versus water-quenched slags. Three slags containing 1.6, 5.0, and 6.6 percent copper were used. More than 90 percent of the copper was recovered in a rougher concentrate leaving a 0.2- to 0.3-percent copper tailings when treating slow-cooled slag. Lower recoveries and higher copper tailings ranging from 0,5 to 0.6 percent copper were obtained from quenched slag. Grindability studies were made on the respective heat-treated slags. Quenched slags proved more difficult to grind than slow-cooled slags. Cost studies showed that (1) the cost of slow cooling and crushing off-sets any advantages of higher copper recoveries from slow-cooled slag, and (2) water-quenched slag, which requires no crushing preparatory to grinding, could be treated more economically. Costs based on processing 1,000 tons of slag per day would be $1.85 per ton for quenched slag versus $2.37 per ton for slow-cooled slag.

Jan 1, 1971

By D. H. Carlson, J. H. Johnson

"The U.S. Bureau of Mines and Michigan Technological University (MTU) are collaborating to develop an apparatus for measuring diesel tailpipe emissions in underground mines. A tailpipe emissions measurement apparatus (EMA) is described that dilutes diesel exhaust and measures the concentrations of diesel particulate matter (DPM), CO, CO2, NO, and NO2 at known dilution ratios.The EMA was evaluated by side-by-side comparison of its measurements with those by laboratory grade instruments. Concentrations of CO and C02 measured by the EMA were usually within ± 14 pct of concentrations measured by laboratory-grade instruments for CO and ±20 pct for C02• EMA measured DPM concentrations were, on average, 29 pct lower than those measured by laboratory-grade instruments, with the differences being fairly consistent. The CO and C02 differences are related to the method of calculating the EMA's dilution ratio, and the difference in DPM concentrations is attributed to thermophoretic and/or condensation losses related to the EMA's unheated sample probe. Preliminary analysis of data obtained with a heated sampling probe and modifications in the sampling procedures have indicated reduced variability in DPM and in the calculated dilution ratios."

Mar 1, 1992

By Paul Soros

Lord Robens, Chairman of the National Coal Board of England announced recently to the world press plans for construction of the largest bulk terminal in Western Europe. This should result in annual freight savings of 22.4 million dollars and between 150 and 190 million dollars of new export business. These are rather substantial amounts for any mining enterprise. Or to put it another way, opportunities to make substantial savings in the delivered price of a highly competitive commodity such as coal are few and far between. One area where such opportunities still exist is in the economical; movement of bulk materials between producer and consumer.

Jan 1, 1966

By Joseph P. Sobanski, Michael P. Bruen

The North Branch tunnel segment of the Tunnel and Reservoir Plan Mainstream System consists of 14,075 m (45,885 ft) of 9.2 m (30.2 ft) finished diameter and985 m (3,232 ft) of 2.5 m (8.25 ft) finished diameter concrete lined rock tunnels. The TBM mining and concrete lining methods used in completing this $166million project differed from “typical” TARP construction methods in that a two-stage concrete lining was placed and the lining was initially placed concurrently with TBM mining as a two pass system in the same tunnel segment. For the main tunnel (9.85 m diameter), TBM mining averaged 24 m/day (79 ft/day), excluding downtime associated with two main bearing changes, and the concrete lining placement, averaged 117 m/day (380 ft/day) and 53 m/day (174 ft/day) for the invert and arch segments, respectively.

Jan 1, 1999

By R Dimitrakopoulos, K Kawahata, L Montiel

A method that optimises mining complexes that are comprised of multiple processing destinations, open pits and underground operations is presented. Mining, blending, processing and transportation decision variables are simultaneously optimised while accounting for geological uncertainty. The method uses a simulated annealing algorithm at different decision levels to generate a stochastic-based extraction sequence and processing policies. A case study shows its ability to generate a higher net present value while facing a reduced amount of risk when compared to traditional optimisation methods.CITATION:Montiel, L, Dimitrakopoulos, R and Kawahata, K, 2014. Globally optimising open pit and underground mining operations under geological uncertainty, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014, pp 201–212 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 24, 2014

By Ch. Enkhbold

Underground mining of minerals is becoming an increasingly complicated technological and ecological problem. It can be explained both by worsening of geological conditions of mining and by a specific character of mining industry, which, in contrast to processing branches of industry, intrudes into one of principal components of the nature ? the Earth's interior. At a traditional organization of mining industry, the entire volume of the produced rock mass is brought to the Earth's surface, where valuable components are extracted. This entails not only power over expenditure for long-distance transfer (especially in the vertical direction) of the waste rock forming a part of the original mineral, but also the formation of huge volumes of worked-out spaces. The latter leads (nature cannot stand emptiness!) to an irreversible development of geomechanical processes of shifting of overlying rock masses resulting in an irreparable damage of all surface structures located on the territory undermined by deep mining. Owing to their high productivity combined with low power consumption and instrumental compactness, novel bifunctional technological processes of minerals dressing successfully solve the problem of rock mass processing in underground conditions with subsequent operative disposal of final tailings of such underground dressing in the worked-out space. Besides, an integration of two processes into one is achieved, for example, a decrease in the rock mass coarseness with a simultaneous precise separation of valuable and waste components.

Aug 1, 2013

By Charles A. Kliche

INTRODUCTION The Northern Great Plains Coal Province which occupies approximately 36.6 ha contains about one-half of the nation's total coal resources. About one million ha are underlain by coal that is considered to be mineable from the surface. The topography of the province is predominately plains, rolling hills, and some badland areas characterized by breaks or sharply eroded hills. Elevations range from about 1500 m above sea level on the west to about 600 m above sea level on the eastern boundary. The average slope of the ground is approximately 2 m per km. Streams in the province drain generally easterly and southeasterly (1). In the United States, the Northern Great Plains Province lies within the states of North Dakota, South Dakota, Montana, Wyoming, Nebraska, and Colorado. The coal within the province is located almost entirely in the states of North Dakota, Montana, and Wyoming. Although it is relatively low rank, the coal within the province encompasses a large portion of the total remaining U. S. reserves. The U. S. Geological Survey estimates the total remaining coal resources of the United States to be 2.6 trillion mt within 1000 m of the surface, and 2.9 trillion tons within 1830 m of the surface. Of this total, only about 177 billion mt are classified as reserves meaning that they are (1) known in location, quantity and quality from geologic evidence supported by engineering measurements, and (2) economically recoverable based on currently available technologies and current market value (2).

Jan 1, 1983

By Diana Drinkwater, Tim Napier-Munn

"This paper argues that, although specialist mineral engineering expertise is essential for the efficient design and operation of mineral processing plants, the environment in which these skills are supplied to the industry has changed forever, and not for the better. Most of the mineral engineering programmes in western universities have died or are dying. The long-service mine site mentors have gone, operations are run on a staffing shoestring so that there is no longer the time (or inclination) for considered decision-making, and the FIFO model has discouraged continuity of optimisation projects and communication between professionals. At the same time the technical challenges of designing and operating effective process plants are increasing. These trends are not going to be reversed any time soon, if ever. The ability to run efficient mineral processing operations has been materially compromised as a consequence. We propose that well-designed professional development can help to recover this situation, and it can be done easily and cheaply. The workplace is the best environment for such skills development. Professional development training should be integrated with normal duties throughout the early years of a graduate’s service (and probably beyond), and the cost should be regarded as non-discretionary expenditure for the business. If such programmes are properly resourced and managed and extended over longer time periods, there will be a measureable rise in the level of technical understanding and competence within companies and a resulting beneficial impact on resource utilisation and thus shareholder value. The paper outlines some of the approaches that we believe will work, and the benefits that we expect will accrue."

Jan 1, 2014

By Alexander N. Winchell

"The deposits of the east end of the Mesabi Range cover very large areas, and consist of the same taconite that produces the hematite of the central and western Mesabi. It is not deeply covered with drift, but outcrops boldly in large masses. It is not concentrated locally into enriched bodies separated by leaner areas, but is all a fairly uniform, hard, banded, unaltered chert, carrying about 25 to 30 per cent iron as magnetite.The Mesabi Iron Company, organized by Hayden, Stone & Co., and under the direction of D. C. Jackling, has taken over these deposits and is building a Concentrating plant at the new town of Babbitt, about fourteen miles east of Mesaba, Minn. This plant is to have a capacity of 2,000 to 3,000 tons per day, and is based on the application to this low grade material of the methods which have made the so-called porphyry copper properties so successful.In 1916, an experimental mill was built in Duluth; containing full sized machinery and capable of handling from 100 to 200 tons daily. A considerable tonnage of material was concentrated here during 1916, 1917 and 1918. The work showed that the resulting product could be varied, almost at will, between 60 per cent and 70 per cent iron and .006 per cent to .030 per' cent phosphorus, and that this could be made on a large scale well within the limits of cost set by iron trade conditions. A small special cargo was sent down the lakes, late in 1918, assaying 63.27 per cent iron (natural and dry) and .008 per cent phosphorus; but it is assumed that large scale shipments will normally be of a Bessemer' grade carrying about 63 per cent iron and .020 per cent to .030 per cent phosphorus. This material is a sinter, carrying only traces of sulphur and titanium, with silica between 6 per cent and 10 , per cent, and alumina, lime, magnesia, and manganese, each below 1 per cent. Some of the iron in this sinter is hematite and some is magnetite, the percentages varying according to the treatment given on the sintering machine. Like the analyses, this hematite percentage is within reasonable control. Sinter made from finely ground magnetite has much the structure of coke, being firm yet porous, and gives good account of itself in the blast furnace."

Jan 1, 1920

By Nathaniel Arbiter

The development of flotation machines since the first industrial use of the process can be divided into three stages. The first, up to about 1930, was marked by diversity in the means for contacting floatable mineral with the air/ water interface and for generating the interface, as well as in cell design. By 1927 Taggart (1) could list 170 patents covering flotation cells, and he mentioned dozens which were or had been in operation. He divided cells into eight categories according to design, type of separating interface, and where in the cell and how separation took place. These were: (1) Film or skin flotation with wet or dry oiling of surfaces; (2) pulp body process with bubbles generated chemically by pressure reduction, by vacuum or by boiling; (3) agitation-froth processes involving local supersaturation with air; (4) bubble column; (5) pneumatic; (6) cascade; (7) subaeration; and (8) combination and miscellaneous types and processes. Most widely used then were the agitation-froth, sub- aeration, pneumatic types. By 1943 a significant reduction in categories had occurred to the two main divisions of mechanical and pneumatic cells according to whether agitation and aeration were (a) by an impeller with or without compressed air, or (b) by compressed air alone. Cells at that time were still relatively small; the Morenci Concentrator completed in 1942 used 432 1.7m3 (60 cu ft) Fagergrens for roughing 45,000 and later up to 55,000 t/d.

Jan 1, 1979

By Christian Bourcier

La méthode d?exploitation traditionnelle de la mine est le chantier chambre-magasin. Cependant, lors de l?exploitation d?un chantier, nous avons été confrontés à un problème de soutirage. Le minerai abattu est resté accroché à l?intérieur du chantier ce qui nous obligea à soutirer environ 200 wagons, représentant 700 tonnes. Suite à l?ampleur de la situation, nous avons décidé de finir l?exploitation du chantier par la méthode sous-niveau long trou retraité. Cette première expérience de récupération de piliers ayant été réalisée avec succès, nous avons décidé de continuer avec deux autres récupérations de piliers afin de pouvoir augmenter notre expertise autant au niveau technique qu?au niveau économique.

Oct 1, 2001

By G W. Wilson

Net solar radiation is one of the key factors affecting evaporation from soil covers on waste rock dumps. While net solar radiation is normally taken as a uniform parameter over a site for modelling purposes, there can be significant variation in the radiation actually received on various portions of the dump. These variations are a function of the difference in the cover slope and direction of exposure. A predictive model has been developed to calculate the actual radiation received on the dump surfaces, based on the net radiation data measured at a weather station on a horizontal surface. The model has been designed for application at any mine site in the world, and can be used in conjunction with evaporative flux models to help estimate the effect of three-dimensional variations in the cover surface exposure. The waste rock dump at Equity Silver site (located in British Columbia, Canada) has a barrier-type soil cover system to limit oxygen entry and water infiltration. This cover has been monitored for the past nine years. The new net solar radiation model has been applied to this site and the predicted net radiation was compared to the values of actual net radiation measured on the site. The predicted variations in net radiation over the three-dimensional surface of the waste rock dump have been applied to calculate spatial variations in potential and actual evaporation (or evapotranspiration) from the site. Preliminary analyses indicate actual evaporation from a site at this latitude could vary by a factor of two for exposed versus shaded slopes. The implications of these variations for the site are considered.

Jan 1, 2003

By Daniel H. White

Block caving, unlike most other conventional forms of mining relies principally on the natural forces within the rock mass to fragment and transport the ore. Obviously, not every orebody can be satisfactorily mined by block caving. In the early stages of mine planning an attempt should be made to assess the benefits of the block caving system relative to others, and develop a reasonable mining concept that may be used to provide preliminary information to mine planners and financial specialists. It is the intention of this paper to outline some basic geotechnical parameters that are essential to: 1) Assessing the feasibility of block caving an orebody and 2) Providing semidetailed information relative to mine design based on drill hole and other data.

Jan 1, 1979

By M. H. Pelkki, R. J. Sweigard, P. W. Conrad, J. M. Ringe, D. H. Graves

Research has shown that excessive soil compaction is detrimental to the establishment of trees in reclaimed soils. Dry bulk density and penetration resistance are two soil parameters that can be used to indicate the level of compaction in reclaimed soils. It is believed that reclamation practices which minimize soil compaction will produce better results when forestland is the intended post-mining use. The University of Kentucky has developed reforestation test cells on reclaimed soils at a surface mine in eastern Kentucky to test this hypothesis. Each test cell has been planted with various species of trees, and since 1997, data have been collected at these test cells for soil average dry bulk density, average penetration resistance, average depth to refusal, and average survival rate of trees. These data have been correlated to determine the relationship between soil compaction and tree survival rate. It is believed that the two soil compaction parameters may ultimately be used to predict tree survival rates on reclaimed surface mined lands.

Jan 1, 2001

By R. V. Ramani, C. B. Manula, A. Owili-Eger

ANALYTICAL PRESSURE - VOLUME COMPUTATIONS FOR THE MINE PROTOTYPE A new coal mine is projected, with entrance by slope 1,500 feet long. Nine main entries will be standard for development, five as intake and four as return. One return shaft is planned near the slope bottom, and an intake shaft and portal are planned 5,000 feet north of the slope bottom. The slope will have an area of 150 square feet and a perimeter of 55 feet. A belt and supply track will be within the slope. The return shaft will be circular with a diameter of 18 feet and will be 500 feet deep with an escapeway. Two hundred feet of the shaft and 500 feet of the slope will be concrete lined. Select the fan to operate five units per shift and drive the 5,000 feet to a proposed new portal, using the slope for an intake. Two main splits are planned with the maximum of three units on any split. The amount of air desired at each unit is 25,000 c.f.m; a shop at the slope bottom will require 10,000 c.f.m. In the solution of this or any similar problem, the first step is to determine the resistance factors. CALCULATION OF R FACTORS Slope area = 150 sq. ft.; Perimeter = 55 feet; Length = 1,500 feet. The friction factor K must be established. Five hundred feet of the slope is smooth lined, moderately obstructed by belt; K = 35. The remaining 1000 feet is in sedimentary rock and moderately obstructed; K = 85.

Jan 1, 1975

By Thomas P. Goslin

T e mechanization of industry to reduce labor costs and to improve product quality is affecting he scale industry as it has affected many other industries. With increased emphasis on higher production rates many operators are experiencing difficulty in overcoming production problems where weight is a factor. Unfortunately, the scale industry has not shown the development over the last twenty-five years that other instrument industries have. For centuries the balance scale has been the standard and the workhorse of most weighing operations. In recent years considerable effort has been applied to designs to make balance scales more acceptable to industries employing automatic or semi-automatic processes. For the most part the designers have not attempted to replace the basic scale but toautomate functions of material handling and human judgement. Modern dump scales, for example, will accept material from a conveyor until a preset amount of material is on the scale, At this time the scale turns off the conveyor, dumps the load of material and repeats the operation with any desired timing. These improvements have helped industry. Even the scale observer has been replaced with devices which observe the scale reading and operate on weighed objects according to their weights. In spite of' the application of automation to the material handling, the balance scale has not been completely satisfactory. One reason for this is fairly evident when one considers the construction of the scales. Inherently, they are slow devices because of the inertia of the beams. The beam scale, primarily a static measuring device, is at the best limited to slow dynamic measurements.

Jan 1, 1958

By V. E. Ross

A comparison was made of three methods for determining the contributions of true flotation and entrainment during batch flotation tests. In the first method size-by-size recoveries are measured in both the presence and the absence of a collector, a frother being present in both tests. In the second method the mass of solids and water recovered after a chosen time of flotation is varied by changing the rate of froth removal and the depth of the froth. The third method uses a transfer factor to describe the classification of solids and water, and hence the changes in the composition of the pulp, during a single flotation test. Batch flotation tests were carried out with artificial mixtures of pyrite and gangue, the amount of floated material recovered being calculated according to each of the three methods. The discrepancies betweeen the calculated amounts are analysed in terms of the assumptions that are made for the various methods and the most suitable conditions for the application of each method are identified

Dec 1, 1991

By J. H. Swisher

ALUMINUM deoxidation equilibrium in liquid iron has been the subject of many investigations. Sawamura and Sano1 have written a critical survey of the literature on this subject and consider the data of Gokcen and chipman2 and of d'Entremont, Guernsey, and Chip-man3 to be most reliable. The present study was undertaken to check these data and extend the range of the experiments to lower temperature. The experiments consisted of equilibrating 5-lb Fe-Al melts with recrystallized alumina crucibles at 1580o + 10oC, and analyzing the solidified ingots for aluminum and oxygen. The electrolytic iron stock contained 100 ppm C, 20 ppm 0, and trace amounts of other elements.* The stock was induction-melted in an atmosphere of 5 pet H2 in argon to attain equilibrium between the melt and the crucible. The temperature was measured with a W, 3 pet Re/W, 26 pet Re thermocouple, and the equilibration time was about 4 hr. The results are plotted in Fig. 1 along with curves obtained from the data of d'Entremont et al.,3 for temperatures of 1740o and 1910°C. The present results can be seen to be consistent with the previously reported high-temperature data. The reaction under study is AI2O3 (s) = 2 Al (l) +3O [1] for which the deoxidation product is Kl =[% Al]2 [%O]3 [2] (Under the experimental conditions, Fe A12O4 is not expected to form.4) In Fig. 2, values of log Ki evaluated at 0.01 pet A1 are plotted against 1/T, using data obtained from all the investigations under consideration. The deoxidation product is plotted in Fig. 2 in lieu of the equilibrium constant (normally obtained by evaluating Ki as pet A1 — 0) because of the uncertainty involved in extrapolating to zero concentration. The difficulty in extrapolating the present data for 1580°C is apparent curvature in the log Ki vs pet Al plot. This curvature is perhaps due to the uncertainty in oxygen analysis for concentrations of 5 ppm and below. For the data of Gokeen and chipman,' the extrapolation is also difficult because of experimental scatter. The line drawn through the points plotted in Fig. 2 may be represented by the equation:

Jan 1, 1968

By S. Mohanan, C. R. Kumar, S. K. Bhoja, T. Venugopalan

"A pressing issue that has often been the topic of discussion in iron-making operations in India is the high content of alumina in the country’s iron ore deposits. The presence of alumina has hampered production by affecting slag fluidity, and thereby the hot metal quality. Iron ores from Indian mines are generally rich in hematite, with traces or small amounts of goethite/limonite, kaolinite, gibbsite and quartz reported from time to time. In day-to-day production, emphasis is placed on checking the chemical analysis of product samples, but the true solution to the alumina problem lies in estimating the mineralogy of the ore samples. We developed a mathematical model, based on singular value decomposition, to predict mineralogy from classical chemical analysis. We then used X-ray diffraction patterns, scanning electron microscope analysis, and grain counting with an optical microscope to determine actual mineral compositions and to test the values predicted by the model. The results indicate that with regular calibration, the model can be used as a simple, fast and effective tool to predict on a regular basis the mineralogy of multiple samples.IntroductionIndian blast furnaces use high-grade iron ores (as lumps or as sinter) and coke (obtained from the destructive distillation of coal) as raw materials for the production of high-quality hot metal to be converted into different steel products through the Linz- Donawitz process. The iron ores are mainly sourced from the Singhbhum region of Odisha and Jharkhand states, the Bailadialla region of Chattisgarh state and the Bellary region of Karnataka and Goa states, and are mainly classified based on size as (1) lump ore (fed directly into blast furnaces), (2) fines (converted into sinter/pellets and fed to blast furnaces) and (3) slimes (< 0.15 mm; currently dumped in huge slime ponds). Any variation in the quality of the ores supplied results in variations in the production trends of hot metal."

Jan 1, 2015

By T. H. Watkins

COAL is the basis upon which the activities of the industrial nations of the world rest today. In order that the coal industry may bee restored to a reasonably healthy condition, it is essential that both the operator and the miner should understand that they are directly concerned and that their selfish interests demand that they take an active, intelligent, self-sacrificing view of the entire situation and do their part in restoring and repairing the economic structure. Foreign credits, foreign debts, unstable exchanges, destroyed markets, and inability through lack of credit to carry on as before, all require our attention and cooperation. DEEP CONCERN AS TO FUTURE There is deep concern as to the future and a realiza-tion on the part of some of these leaders of industry that they will have to sacrifice their personal time, away from their desks and the details of business, in order to cooperate with other groups in studying the situation and endeavoring to find the best method of utilizing their knowledge in improving conditions and averting further disaster. It is most unfortunate for the country and for labor itself that the labor chiefs are blind to these necessities, and hostile to every effort to secure their cooperation in rebuilding prosperity.

Jan 2, 1922