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By R. W. Barbaro, R. V. Ramani

Generally, production planning is concerned with the best use of fixed resources to determine a production schedule for one or more production units supplying one or more markets. Production scheduling determines when each production unit will be used and its production level. However, the production scheduling problems are intimately associated with the location of existing or new production units and processing facilities. Normally, mine location allows only a limited choice; processing plant types and locations are considerably more flexible. This paper presents a generalized multiperiod MIP model as an aid in determining the mine production schedule and the processing plant types and location to satisfy multiple markets. A brief description of MIP and how it has been applied in the mineral and other industries is discussed. A complete description of the model with an application to a coal preparation plant system is presented.

Jan 1, 1987

By C. D. Wittreich

The results of an engineering study of fiber-reinforced cemented tailings are presented. These tests were conducted to evaluate the potential for using fibers to increase the strength of cemented tailings used as backfill for ground support in the undercut -and - fill mining method in a mine subject to high horizontal stresses and rock bursts. Stress-strain measurements indicate that fibers not only increase the peak compressive tensile and flexural strengths of cemented tailings, but limit reductions in residual strength following failure. Results also show the effect of water content, fiber type, fiber density, and curing time on strength.

Jan 1, 1987

By D. G. Krige

The problem addressed is that of estimating grade differences between sub- divisions of a new gold mining prospect when only borehole results are available. Such differential grade estimates are required in choosing the best area for initial production, and in defining the grade profile over the mine's life for purposes of feasibility studies. It is also often necessary to differentiate between the grades and hence the values of subareas, owned by separate companies, for the purpose of fixing relative share participation percentages in the new mining company. It is shown that exaggerated grade differentials will be obtained from grade estimates of the subareas based only on boreholes within each such subarea, respectively, and that it would be more realistic to apply some 'smoothing', preferably by kriging. This is feasible because the lognormal-de Wijsian model can be used to give an indication of the spatial structure for the boreholes, even where they are spaced far apart; also the degree of 'smoothing' indicated is shown to be relatively insensitive to inaccurate estimates of the relevant semivariogram parameters.

Jan 1, 1987

By Carla A. Kertis, Alan A. Campoli, Claude A. Goode

This Bureau of Mines study was conducted to obtain a better understanding of the coal mine bump problem and its effect on underground coal mining in the Eastern United States. To accomplish this, information was collected on the geologic conditions, mining techniques, and engineering parameters at five bump-prone mines. Two geologic conditions have been found to cause the occurrence of bumps in the Eastern United States: (1) relatively thick overburden and (2) extremely rigid strata occurring immediately above and below the mine coalbed. Additionally, the probability of bump occurrence is increased by certain mining practices that concentrate stresses during retreat mining, in areas where geologic conditions are conducive to bumps. Mining plans that permit the development of pillar line points or long roof spans that project over gob areas should be avoided because these features may contribute to the occurrence of bumps.

Jan 1, 1987

By Yu-Jun Z

According to site measurement and labora- tory experiment, the authors suggest a new mathematical model of grouted rock bolts. The rock bolt is taken as an elastic bar, and the grout is substituted by a set of springs the rigidity of which is given as K. Based on certain assumptions, a simplified formula is deduced. When the category of rock, depth of tunnel, the spacing of rock bolts and the rigidity of springs K values are given, so that we can calculate the radius of plastic zone, and subsequently the convergence of rock at the tunnel inner surface u, and the resis- tance exerted by the rock bolt p. Finally, the length of rock bolt can be determined.

Jan 1, 1987

By Joseph M. Denk, John N. Baran, William J. Francart

Filling or sealing shafts of underground coal mines is required when the mine is declared inactive by the operator, permanently closed or abandoned for a period of time greater than 90 days. Although shaft filling requires the use of incombustible materials for the first fifty feet of the shaft, a sparking potential may exist which has been reported to be the cause in at least one shaft-filling explosion. This paper outlines the method used at the explosion site to monitor mine gases within the shaft to assure the safety of the workers involved in the remainder of the shaft-filling operations. Gases which were monitored include methane, oxygen, carbon monoxide and carbon dioxide. Continuous sampling was conducted twenty-four hours per day to determine the methane liberation and accumulation of gases within the shaft. Once gas concentrations were evaluated, the need for ventilation in the shaft was obvious because methane accumulations within the shaft were in the explosive range. Several methods to ventilate the shaft were explored. The most practical seemed to be the dilution of gas concentrations with compressed air to inject a quantity of fresh air into the shaft. But, when the compressed air was stopped, a buildup of an explosive mixture reoccurred. The pattern of gas accumulation was then studied to find the correct combination of ventilation and shaft-filling time periods which would a1 low for the greatest reduction of the potential for an explosion. Gas concentration sampling continued during the remainder of shaft-filling operations at this shaft and four others at the same mine. In conclusion, this paper illustrates some of the problems which are often overlooked in a shaft-filling operation. If addressed, the work can be done in a safe and feasible manner.

Jan 1, 1987

By Harold D. Stevenson, Gerald G. Carlson

The Flow-Through Shares Program was set up by the Canadian Federal Government in 1978. While initial annual expenditures in flow-through shares was small, recognition of flow-through as a significant tax advantage combined with exploration successes, particularly in the gold sector, soon resulted in the majority of exploration dollars raised in Canada utilizing the flow-through share mechanism. What Are Flow-Through Shares As many exploration companies are non-taxable, they are not in a position to benefit from the tax deductions arising out of their exploration activities. Flow-through financing provides a mechanism for the exploration company to transfer these deductions to an investor. In addition to the deductions which an investor can apply against his personal income, the investor receives common shares in the exploration company (flow-through shares). To add to the attraction, the Federal government provides a 33 1/3 percent mark-up on qualifying expenses incurred by the exploration company, giving the investor a deduction of up to $1.33 for every $1.00 invested. By providing the investor with tax deductions in addition to the flow-through shares, flow through shares can often be sold at a premium to market which makes-the mechanism attractive to the exploration company as well. These shares are often sold through private placements and therefore have to be held for one year. The tax cost of the share is deemed to be ril, so the entire proceeds on disposal are

Jan 1, 1987

By R. P. King

A class of gravity separation unit operations in ore-dressing relies for its operation on the production of a stratified bed. Heavier material gravitates downwards, and separation is achieved by splitting the ore horizontally, with the recovery of a dense stream and a lighter stream. The important ore-dressing unit operations that fall in this class are the jig, the Reichert cone and the pinched sluice. A new quantitative model based on Mayer's potential theory is developed here. This model is developed completely for the simple case of binary mono-size particulate materials, but the extension to units operating with feed material distributed over particle size and density is discussed in principle. The model can be used to predict the performance of these units and is very suitable for simulation. The model is tested successfully against experimental data obtained from a batch jig and a Reichert cone.

Jan 1, 1987

By M. Gershon

The article is an excellent one, providing a basis for the conduct of the major tasks in underground mine planning and scheduling. I object to only one word. Since that word is featured in the title, it is a strong objection. The word used erroneously is optimization. There is no optimization performed in the article. By definition, an optimization is an approach that yields the best of all possible solutions. We can limit this definition to its narrowest sense, the optimization techniques of operations research such as linear programming, or broaden it to include less sophisticated search procedures that find the best mine plan or schedule. No such claim is made in the article. No optimization of any kind is performed. The major tool used is a simulation, which, in truth, is probably more appropriate for this problem than any optimization. But simulation does not optimize. It only allows for quick evaluation of many options, leading to the selection of the best plan "of those tried," not "of all possible plans." This article, of course, is not the first mining paper to make this error. It seems that the word optimization is used more to impress people than to refer to any act of optimizing. It is used in a way that implies "try to optimize" or "come close to optimize," but not to really do it. It is usually appropriate to quietly ignore this error. But when Dr. Chatterjee, who has done so much good work in true optimization projects and whose work I have come to admire, makes such an error, it is time to publicly dispute it. Having discussed this very point with him in the past, I am sure that he will agree. If he can support this position in his response, perhaps people will start to listen and we can reduce the number of false claims of optimization. This should be easy for him in this case since the approach he used, simulation, is the more appropriate one anyway.

Jan 1, 1987

By M. Bailey-McEwan

The application of refrigeration on deep South African gold mines is mostly achieved through the medium of chilled water which is produced in central installations either on surface or underground. The amount of refrigeration has been increasing rapidly over recent years, and the total installed capacity on all gold mines now exceeds 1000 MW(R), with annual running costs exceeding R100 million for electricity alone. Hence, there are strong incentives to design such installations for optimal cost-effectiveness. The duty required of these installations varies considerably, owing to daily and seasonal climatic changes and the varying nature of the water demand. It has long been recognized that a computer program to predict installation performance as a whole under any given conditions would be highly desirable. An interactive computer program, CHILLER, is described which enables the user to build an installation, consisting of any configuration of cooling towers, water storage reservoirs and packaged water chilling machines. Externally-dictated conditions such as the weather and chilled water demand must be specified, and all quantities under control of the installation, including internal water flow rates and compressor capacity regulators, must be set. The time-varying performance of the installation can then be simulated. Examples are given of the application of the program in the design of a new installation, and in evaluating the effect of any modifications to existing installations.

Jan 1, 1987

By D. A. Huggins

Column flotation is finding increasing use in the flotation of copper, molybdenum and other heavy metals and coal. Columns have proven operational, cost and metallurgical advantages which caused us to investigate their use in the proposed Escondida flowsheet. A comparative pilot plant test with 35 tons of ore showed that one column can replace three stages of cleaning and result in a significantly higher concentrate grade with no loss in recovery.

Jan 1, 1987

By T. Hackett

Western underground coal reserves are tremendous, but recovery in western seams can be poor because of mining problems caused by deep cover, geologic anomalies, and thick seams. For instance, many seams contain a rock parting too thick to be extracted economically with the coal. This forces the operator to mine the coal above or below the parting. Subsequently, the rock pressures induced by mining make the part left difficult or impossible to extract by standard mining methods. Because upper and lower parts of the seam are frequently of minable thickness, the presence of rock partings can result in a significant loss of coal. Given a longwall recovery of 60 pct and equal thickness for the two parts of the seam, overall recovery is reduced to less than 30 pct if one part only is mined. The coal left in place is unavailable to meet the nation's energy needs, or to produce revenues for coal companies or government. [ ] Multi-slice longwall mining has the potential to recover the portion of split seams that are now left in the mine. Using superimposed standard longwalls with chain pillars, overall recovery could be doubled, depending on panel and pillar widths. Using a combination of advancing and retreating longwalls in descending slices, overall recovery could be improved to more than 80 pct. Because of the potential for improved recovery, the Bureau of Mines is analyzing the ground control factors affecting multi-slice longwall under western mining conditions and is conducting field trials at a planned multi-slice operation.

Jan 1, 1987

By R. P. Vinson

The Bureau of Mines evaluated two real-time light-scattering dust monitors by measuring their response to Pittsburgh Seam coal dust and Arizona road dust (ARD). Both monitors, the model PDS-l, a personal dust sensor, and the model PCD-l, a digital dust indicator, are made in Japan by Sibata Scientific Ltd. Tests were conducted inside a laboratory aerosol chamber designed to maintain a uniform spatial distribution of the test dust. PDS-l and PCD-l measurements of concentrations for each test dust were averaged over 4-h test periods and compared with simultaneous gravimetric measurements. Both dust monitors responded linearly with mass concentration with both coal dust and ARD. However, the linear response of the PDS-l differed from that of the PCD-l for both dusts.

Jan 1, 1987

Pulverised coal firing was pioneered in the cement industry prior to the turn of the century. Ash from the coal is absorbed into the mix in the kiln and replaces secondary materials. With high purity limestone coal with high ash content can be used. Fuel cost is a major part of the total manufacturing cost. Cement plants are thus quick to respond to fuel price changes and will use low grade fuels if there is a price advantage. Combustion conditions in a cement kiln differ markedly from a boiler. Flame length is determined by air/fuel mixing but because of fixed secondary air conditions this depends on primary air momentum. Volatile content has some effect on flame length but this can be offset by coal fineness. Much of the heat for cement production is for limestone calcination which is a low grade energy requirement. By allowing long residence time in separate direct fired vessels which then feed the high temperature kiln the energy of materials such as oil shale, tyres and lump lignite can be utilised.

Jan 1, 1987

By S. G. Malghan, W. H. Moyer

In reference to the section "Excess pyrite," the author discusses difficulties in separating chalcopyrite from pyrite. A technique does exist, and has been used commercially, for the flotation, separation, and recovery, if desired, of both chalcopyrite and pyrite. The method was described in Mining Engineering, March 1966, "An Improved Flotation Method for Cornwall Sulfides." While it was not used on a porphyry copper ore, it was used on a chalcopyrite- pyrite-containing ore and might well have application. Both sulfides, chalcopyrite and pyrite, were floated simultaneously using xanthate as a collector and MIBC frother to produce a cleaned sulfide product. This concentrate was then thickened and conditioned at high solids with aeration and lime to a pH above 12.0 for a period of 1.5 to 2 hr. A supplemental suppressant, such as cyanide, could be used if necessary to sharpen the subsequent separation; however, it was not required at Cornwall. The conditioned sulfides were floated with two stages of cleaning to produce a high- grade chalcopyrite froth product and a pyrite sink product barren in copper. The separation at Cornwall gave a 28.6% Cu grade chalcopyrite product and a pyrite analyzing 0.06% Cu. A desliming step on the raw ore initially used at Cornwall for a period of time was later discontinued. Advantages to be found in using a strong sulfide collector, such as a xanthate, is the rapid flotation it induces and, generally, an ability to float at a coarser grind. A following regrind circuit could be used to give maximum liberation and produce a clean sulfide concentration. The "differential" separation is concerned primarily with grade and with assuring a barren pyrite for maximum recovery.

Jan 1, 1987

By Hans Hamrin

Three drill-blast cycles of 3.1 m advance achieved in one working shift of 7.5 hours at the Holsfjorden Tunnel project in Norway. The tunnel, 5235 m long, with a cross section of 8.6 m2, was completed several months ahead of schedule. This remarkable performance was achieved by a crew of 16 men, using machines specially designed for small tunnel excavation. The integration of drilling and charging into one unit, mucking and transport in another, reduced non-productive time considerably. The average cycle time of 140 min made it possible to complete three rounds in one 7.5 hours shift.

Jan 1, 1987

The Proterozoic Pb-Zn deposits of the Gulf of Carpentaria, Australia, include HYC (McArthur River), Mt. Isa, Dugald River, Lady Loretta and Hilton. Two seamingly incom- patible depositional models have been proposed to account for their origin. Finlow-Bates (1980), Large (1983) and Russell(1983) classify these-deposits as sediment-hosted submarine exhalative deposits. This class includes Rammelsberg, Meggen, Silvermines, Sullivan and others. In contrast, a lacus- trine, shallow-water origin has been advocated by Williams and Logan (1981), Logan and Williams (1984), Muir (1983) and Eugster (1985). The discrepancies between these two models can be removed by creating a separate class, that of continental sediment-hosted exhalative deposits.. HYC., Mt. Isa and Dugald belong to this class and their sediments were deposited in shallow, alkaline lakes. Metal- rich, hot, Na-Ca-Cl brines entered the basins along boundary faults and interacted with pore fluids and sediments, depositing-Cu-Zn-Pb sul- fides. Similar brines entered the Lady Loretta basin after marine flooding, depositing sulfi- des and barite, placing Lady Loretta into the class of marine Sedex deposits.

Jan 1, 1987

Brief summary of results obtained by a French-Japanese cooperative study on the bottom of trenches around Japan. The geophysical prospection, multi beam echo sounding and direct observation with submersible dives of the trench bottom clarified the details of the subduction of oceanic plates. It may be noted, among many new facts found: 1. The subduction and geologic structure resulted from the process is different according to the nature of subducting oceanic plate, and the geology of the land behind the trench. Thus the Nankai trough characterized by the developpement of accretionary prisms is contrasted to that of the Japan Trench where the basement rocks are exposed on the land side cliff of the trench wall. 2. The compressional tectonic situation prevailing in the trench bottom provokes the venting, along the fissures on the bottom, of water contained in sediments. Hydrocarbons contained in the sediments are comming out with water and give nutrients for bacteria. Clams, tube warms, sea slugs and shrimps are taking these bacteria as their food and biological communities with exceptionally dense population for the deep sea of more than 3000 meters in depth are formed on the field of accretionary prism and wall of the land side cliff. 3. An intense compressional tectonic field resulted from the subduction is inducing the transition from the subduction of oceanic plate to the collision of a part of the plate to the continental plate.

Jan 1, 1987

By T. A. Edison

A distinct hydrocarbon maturation anomaly coincides with the Gold Acres open pit, Nevada (Figure 1). Both atomic hydrogen to carbon (H-C) ratios and reflectance measurements of contained organic matter indicate significantly higher maturity in samples collected within the confines of the open pit than in samples from the surrounding region (not shown in Figure 1).

Jan 1, 1987

The accuracy of an ore reserve is only as good as the geologist's understanding of the deposit. The methods that are applied to resource and reserve estimation reflect that understanding. At the Olympic Dam copper-uranium-gold deposit in South Australia, evolution in the understanding of the controls on mineralisation coupled with the changing demands of the project have led to changes in the approach to reserve estimation. The project has moved into a phase where detailed stope mining reserves are now required as distinct from global ore reserves. The complex nature of the mineralisation and the large amount of data demand that all available information be readily accessible in a flexible format. To enable the selective manipulation of geological and assay information and its characterisation, a relational database has been developed. For reserve calculations themselves, initial computations were based on a system derived from that used for the Kambalda nickel orebodies. The Olympic Dam system differed mainly in the use of statistical analyses in the estimation of grade instead of the previous polygonal area of influence weighting method, Three dimensional weighting techniques are now being used for local reserve estimates. Flexible facilities for raw data handling combined with grade estimation techniques which are commensurate with the mine geologist's understanding, produce detailed gridded and contoured ore zone models suitable for reserve calculations and stope designs.

Jan 1, 1987