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By Harold J. Miller

Loading and haulage of run-of-mine ore is frequently the highest cost in underground mining operations. These costs can be significantly reduced by use of belt conveyors, a common practice in coal and similar soft miner ii mines. This is particularly true if oversize ore and rock can he eliminated prior to conveying; thus producing a demand for crushers near the production areas. Portable feeder-breakers are extensively used in conveyorized mines, but no suitable, equivalent equipment, is available to provide these benefits to hard rock mines of low to medium headroom. A hard rock crusher capable of easy movement and readily disassembled would allow use to belt conveyors and could improve the general safety and productivity of any underground ore hauling system. This report summarizes proposals and findings made in conjunction with the newly designed M-44 Low Head Portable Crusher which is believed well capable of meeting the requirements of a "portable, underground, hard rock" machine as previously defined in earlier reports. Data gathered by use of the 1/4 Scale Model of the M-44 has substantiated and verified the concept sufficiently to justify construction of a full-size machine for further testing under actual mine conditions.

Jan 1, 1978

By R. Sweennen, A. Vervoort, J. Caers

A case study is presented of the mechanical behaviour of fractures in the Westphalian A and B siliciclastic succession of mud-, silt- and sandstones with intercalating coal layers in the Campine basin of northeast Belgium. Mining ceased in the area in 1992 and a pilot well was drilled to assess the feasibility of methane extraction by hydraulic fracturing. Work was carried out to study the influence of types of cement fill in fractures on the shear behaviour of the rocks by comparison with that of the coal seams. X-ray computerised tomography was used for visual appraisal and selection of jointed rocks before destructive shear testing was carried out. Two major type of veins were found to be present, one cemented with carbonate and the other with kaolinite or kaolinite/carbonate. Uniaxial compressive strength and shear tests showed that differentiation of the two types of veins was vital in evaluating the strength of the sandstones, as cementing of the fractures by rigid carbonate minerals can reinstate the strength of the rock masses to that of the intact material. Kaolinite was more susceptible to plastic deformation than carbonates

Jun 19, 1905

By M. Radwan, S. Shimada, S. Z. El-Tawil, K. A. El-Barawy, Y. Miyamoto

This study presents the results of synthesizing monolithic siliconoxynitride, Si2N2O, ceramic from a mixture composed of reclaimed silicon and natural desert sand by using the nitriding combustion synthesis method. Single-phase Si2N2O powder was prepared by the combustion of a mixture composed of silicon and pulverized sand in 3 MPa nitrogen gas. The combustion reaction completed in 2 min and the combustion temperature reached 1577°C.The as-synthesized Si2N2O powder is agglomerate of fine particles with sizes < 5 µm and has various morphologies; rod-like and irregular crystals, irregular kinking, hollow crystals and grown pseudo-hexagonal prismatic crystals. Porous Si2N2O compacts (24.5–47.1% porosities) were fabricated from the synthesized Si2N2O powder at different calcia-alumina additives (5–20 wt%) by normal sintering at 1500°C for 1 hr in N2atmosphere.They showed superior oxidation-resistance at high temperatures up to 1500°C in an oxidizing atmosphere of Ar/O2/H2O(70/20/10 kPa).Near-full dense Si2N2O compacts were also successfully fabricated from the synthesized Si2N2O powder with 5 wt% oxide additives by using the spark plasma sintering at 1600°C.DenseSi2N2O had good mechanical properties (18.7 GPa Vickers hardness, 3.3 MPa.m1/2fracture toughness and 363 MPa strength) and excellent chemical and oxidation resistance.

Jan 1, 2004

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 Frank Marotta, Louis B. Lesem, Frank Greytok, John J. McKetta

Although one of the primary variables in the calculation of the flowing bottom-hole pressure in gas wells from surface measurements is the temperature at any point and its distribution in the flow-string, only few experimental data are available in the literature and little attention has been given to analysis of the problem. Virtually all of the recently published methods of calculating flowing bottom-hole pressures depend on the assumption that either the temperature is constant at some average value or that the variation is linear with depth. The purposes of this work are to analyze the problem theoretically and to verify the analysis by comparison with experimental data so that practical problems in the analysis of the behavior of gas wells can be solved with greater accuracy, reliability, and ease. MATHEMATICAL OUTLINE OF PROBLEM Assuming that: 1. The mass velocity and chemical composition of the gas stream are constant and in normal gas well operations, the change in linear velocity in the entire flow-string is trivial, 2. The product of the density and heat capacity of the gas is constant, 3. No horizontal temperature gradient exists in the gas stream, 4. Net flow of heat by conduction within the formation and in the gas stream in the vertical direction is trivial in magnitude and can be neglected, 5. The regional vertical geothermal gradient is constant, and 6. The temperature of the gas entering the borehole is constant, and equal to that of the reservoir, a set of two simultaneous, linear partial differential equations with appropriate boundary conditions was derived to describe the temperature distribution in the gas stream and the surrounding formation. These equations were solved by operational techniques for the distribution of temperature in the gas stream. The resulting integrals were evaluated numerically on an IBM 604 at the Machine Accounting Div. of the Railroad Commission of Texas. DERIVATION OF FUNCTIONS FOR THE DISTRIBUTION OF TEMPERATURE The physical system is a circular hole of radius a in an infinite medium with thermal conductivity K, density p8, and heat capacity c,. Under shut-in conditions the well is in thermal equilibrium with the surroundings and the increment in temperature is given by - ?T3/L x. Gas of density pg and heat capacity c.. flows upwardly through this hole in the direction of increasing x with a linear velocity W. Writing a heat balance on an element of gas ?x in thickness, convective heat transfer + conductive heat transfer + energy required to lift a unit mass of fluid = change in energy content. pa2W po Co To - pa2Wpo Co To x+xx - 2apK?x ?T8/?r r=a - Wpa2 po?x/778 = CoPopa2?x?To/?t... Dividing by (copopa2?x) and taking the limit as ?x tends to zero, -W (?To/?x + 1/778Co) + 2 K/a Po Co ?T3/?r r=u = ?To/?t

Jan 1, 1958

By D. Large, M. H. Davies, B. Driedrichs, U. Matzander

Synthesis report on a project carried out under the EC-funded BRITE-EURAM programme. A prototype high-resolution electromagnetic system, based on the theory of controlled-source magnetotellurics, was developed. The electrical source field is fed into the earth by dipoles at preselected frequencies and a multi-channel receiver measures five electric and three orthogonal magnetic fields. Suitable mapping algorithms and spatial filters are applied to the data and a new technique permits calibration during measurement. The system was field tested over a blind orebody, at 250-550 m depth, in the Aguas Tenidas area of the Iberian Pyrite Belt. Ten lines totalling 12 000 m of profile length were completed in seven days and the orebody was identified. Tests in the Irish Midlands suffered from strong electromagnetic interference of cultural origin

Jun 21, 1905

By D Will

RedrossThe Redross orebody averages 1 m wide, has a strike of 50 - 350 m, dips at 55° and has a basalt footwall and an ultramafic hanging wall. The most successful and innovative aspects of mining Redross were:•ore development innovations, 90 per cent of development was split fired with a shanty back profile to enable larger and more productive equipment to be used and resulted in higher development grades, improved project payback time and improved ore recovery;•hand held stoping using a multiple level, bottom up ‘V’ shaped hand held stoping method that allowed mining on multiple levels with hand held productivities between 50 and 90 tonnes per shift;•the use of 1 m cement pillars and 7 m high crown pillars were used to create multiple mining areas that enabled high overall production rates from hand held mining this orebody; and•the use of vertical cement pillars to enable hand held mining of adjacent stopes without leaving ore pillars.MarinersThe Mariners orebody averages 3 m wide, has a strike length of 50 - 250 m and dips at 50° with a basalt footwall and poor ultramafic hanging wall. The most successful innovative ideas of re-opening Mariners were:•Dewatering the existing mine workings utilising ‘oilfield’ technology by installing a long submersible pump in on existing pump rising main. This enabled the mine to be dewatered at rates up to 100l/ sec and helped overcome the 45 to 55 L/s inflows.•Overcoming the technical problems associated with Mining N07 ore blocks which had severely deteriorated from being underwater for five years by using reverse firing techniques to improve extraction rates and reduce dilution.•Mass firing a large mining block in the N08 with minimal void space because of deteriorating ground conditions.•Use of a longhole version of the Redross hand held mining method to achieve 100 per cent extraction of the N09 higher grade lenses.

Mar 21, 2011

By M. Kumral

The efficiency of an ore-processing unit depends on the consistency of the characteristics of raw material entering the plant. When the mined ore is highly variable in quality, the only way to ensure consistency is to homogenize the ore prior to feeding to the processing plant. The homogenization can generally be achieved by the bed blending operation. Given that the stockpiling and reclamation processes are very expensive, it is necessary to design the process in such a way as to minimize variabilities of specified properties of raw material. In this paper, for alternative stacking types, optimal stockpile geometry is found in three stages: First, stockpile input is simulated by sequential Gaussian simulation, and then the variance reduction ratios (VRR) as a criterion of stockpile efficiency are calculated for various stockpile geometry scenarios by a stockpile simulator written in FORTRAN. Second, multiple regression analysis is performed to model the VRR by the use of stockpile length, the number of layers and stacker speed as the independent variables. Finally, the model is an optimization problem. Decision variables are the stockpile length, the number of layers, stacker speed and stacking type. The genetic algorithms (GA) are used to minimize the VRR. The approach was demonstrated on data from an iron orebody. The problem was to reduce fluctuations of iron, silica, alumina and lime contents in the stockpile output. The results showed that the approach could be used for the bed blending design efficiently. Keywords: bed blending design, multiple regression analysis, genetic algorithms, iron ore, content fluctuation

Jan 1, 2006

By Gennaro Gerald Marino

Vast improvement in methods of design and analysis of the long- term stability of overburden above room and pillar mines is needed. Improved methods are vital for planning and land development and for prediction above existing, presently stable, room and pillar mines. As shown on the flow chart in [Figure 1], the assessment of the long-term stability is an integral part of the site investigation process regarding mitigation of surface damage effects. In Illinois, flat farmlands with subjacent coal reserves cover vast areas. In some of these areas high extraction of the underlying coal layer(s) may not be possible since the resulting surf ace subsidence combined with a pre-existing shallow water table would make the land too wet. Thus recovery of this coal might only be accomplished by room and pillar mining where the ground surface does not subside and the hydraulics of the overburden are not significantly affected.

Jan 1, 1986

By E. L. Singleton

The Bureau of Mines investigated extraction of aluminum from aluminum-silicon alloys as a means of utilizing domestic aluminum silicate ores for aluminum production. Aluminum was extracted from aluminum-silicon alloys by electrolysis in molten electrolytes of NaCl-KCI-AIC13 and NaCl-KCI-AIF3 in a laboratory-scale cell. Six alloys with aluminum contents ranging from 20 to 87 percent aluminum were tested, and aluminum products of primary grade or better were obtained. More than 80 percent of the aluminum was recovered with current efficiencies as high as 91 percent. Recovery of a metallurgical-grade silicon product from the anode residue by an acid leaching procedure was possible with the 50-50 and 65-35 Al-Si alloys.

Jan 1, 1972

On the more important mines it is almost essential to have a general plan of the surface. A special feature of this plan should be the indication thereon of a suitable arrangement of contour lines, either wholly over the surface of the property or at parts where may seem most useful. The importance of having these contours marked on the plan will be best appreciated by the engineer, in selecting the best available site for a proposed new plant, in locating a tramway, in connecting the surface with the underground workings to the best advantage, or in solving many similar problems that come within his province. The closer the contours are together vertically the better, the limitation being the trouble and expense of getting them on to the plan. They may range from five to fifty feet apart but a distance of ten feet will be found, in most cases, to be the most useful.In deciding the datum level from which these contours are to be measured, it is best to depart from the ordinary sea level datum, and use instead the highest available point on the property. This datum will be called zero and all levels will be reckoned so many feet below it; thus by a contour marked 160 on the plan, it will be understood that the level of the plane of the contour is 160 feet below the highest available point on the property. This system of measuring levels downward is readily suggested from measuring the depths of shafts from the brace or of excavations from the original surface.

By T. T. Chen

Copper anodes generated from copper scrap are relatively rich in Sn, Zn and Fe. Because of the presence of oxygen, only trace amounts of Sn occur in solid solution in the copper crystals. Thus, Sn occurs mainly as SnO2, Cu-Sn-Ni oxide and Sn-Ni-Zn oxide in the copper anodes, and Zn and Fe occur mainly as minor constituents in the Sn-Ni-Zn oxide and Cu-Sn-Ni oxide. During electrorefining, these oxide species are liberated from the copper anodes and report directly to the anode slimes. The trace amounts of Sn present in solid solution in the copper crystals dissolve but reprecipitate as Sn arsenate or a Sb-As-Sn-O phase. Accordingly, high Sn contents in copper anodes do not lead to elevated Sn concentrations in the electrolyte.

Jan 1, 2010

By Eduardo Lozano, Vilem Petr

This paper provides an overview of a new method for avalanche control using a high explosive charge placed at the avalanche runout zone: the case charge technique. This new avalanche control technique is well-suited to mountain roads presenting several advantages over the existing methods: minimum deployment time, low cost, reduced probability of misfires, and decreased operation time in the avalanche danger zone. The research work presented in this paper was conducted via large scale laboratory experiments where the detonation velocity and the blast physical properties were measured. During the charge design, several parameters were controlled: charge mass, diameter, geometry, density, and external confinement. In addition, the effective range of case charges was increased by the use of the Mach reflection for certain topographical situations. Finally, field test results collected by the Colorado Department of Transportation (CDOT) during the 2017 winter season are presented. The CDOT avalanche control crews have adopted the new case charge as the preferred technique in numerous avalanche locations along the Rocky Mountains roads.

Jan 1, 2018

By Claude Cunningham

The primary reason for blasting is to fragment rock. In production blasting, the fragment sizes produced are known to exercise an overwhelming influence over working costs: handling costs and tonnage throughput are directly affected by secondary breaking, loading and crushing, all functions of fragmentation. In addition, particular size fractions are frequently of intense interest in terms of market or process requirements.

Jan 1, 1995

By J. M. Sands

Production of oil in Oklahoma during 1927 amounted to 273,256,900 bbl. (Table l), an increase of nearly 100,000,000 bbl. over the previous year. All of the major fields declined with the exception of those of the Seminole district which produced 136,105,000 bbl., almost one-half of the total production of the state. Because of this large new production, most of the development in the state during the year took place in that district. The Seminole City pool located in T.-9-N., R.-6-E., was the first of these pools to be developed. The Indian Territory Illuminating Oil Co. brought in the first well in this field in March, 1926, for 1100 bbl. This well is located in Sec. 24-T.-9-N., R-6-E., and produced from the Hunton lime reached at depth of 3900 ft. The first Wilcox production was developed by the Independent Oil & Gas Co. and R. S. Garland in Sec. 26, T.-9-N, R-6-E. This well was being drilled for production in the Hunton lime but was dry in this horizon and was deepened to the Wilcox sand in July, 1926, and at the total depth of 4069 ft. developed a maximum production of 8000 bbl. daily. It held up in wonderful manner and was the cause of not only the development of this pool but the prospecting of the whole district for the Wilcox sand. This led to the discovery of the other major pools in that district. Seminole City pool reached a peak production of 264,500 bbl. on February 22, 1927, and slowly declined so that by December it had a daily average of 56,200 bbl. from 321 wells. The second pool in this district to be discovered was the Searight pool which was discovered by F. J. Searight et al. in Sec. 33, T-10, R-6, in April, 1926. The first well was drilled to the Hunton lime encountered at 4115 ft., and came in for 1300 bbl. daily. During August, 1926, the second well was drilled on the same farm through the Hunton lime to the Wilcox sand which was encountered at 4360 ft. It came in for an initial production of 4000 bbl. daily. This started active drilling in this field for the Wilcox sand. The field had a peak production of 40,621 bbl. on June 15, 1927, and declined to a daily average of 22,300 bbl. from 73 wells during December.

Jan 1, 1928

By E. O. RHODES

Germany's' wartime coal-tar industry had two main branches, which can best be considered individually because their raw materials, operations, and distributions were almost completely separated. One branch, old and well-established, was concerned almost entirely with processing coal tar produced by the high-temperature carbonization of caking bituminous coal. Carbonization was effected principally in byproduct- coke ovens and to less extent in vertical, horizontal, and inclined retorts. The high-temperature tar was all processed in tar-distilling plants that handled high-temperature tar almost exclusively.

Sep 1, 1947

By Leonard Weixler, Peter Halbach, Michael Jarowinsky, Johannes Post

Germany as an industrial nation is highly dependent on the import of mineral or metal resources. Sources which have not been considered until now are the marine mineral resources (MMR), which in recent times are enjoying more and more increasing interests and which might provide in the future a considerable contribution to the supply security on metal resources not only for the industrial countries. Backed up by the German government’s "National Masterplan Maritime Technologies" (NMMT) and corresponding industrial R&D-activities, the idea and the advantages of deep-sea mining is gaining more and more interests and supporters with respect to industrial and political considerations on strategic raw material resources as well as on the development of deep-sea mining technologies. The NMMT aims to initiate growth and increase the competitiveness of the German maritime industry on international markets and covers about 10 different marine and ocean technology sectors including deep-sea mining.

Sep 1, 2014

By H G. Bosmare, G Bedser, C V. B Cunningham

Peak Quany has achieved substantial improvements to its bottom lie by giving close attention to blasting practice, and in pr@icnlar by introducing electronic delay detonators for routine production blasting. The precision and flexibility provided by these units has enabled the quarry to gain control over fmes and curtail oversize without increasing explosive usage. In addition, the extent of nuisance from blasting vibrations has been reduced.

Jan 1, 1998

By Jeffrey P. Schubert

The study site in northeastern Illinois underwent surface mining for coal in 1943 and was reclaimed by the state of Illinois in 1972-1973. Additional soil amendments and seed were applied to the spoil surface in 1975-1980. Previous studies in 1975-1978 had shown that the surface spoil materials and five final-cut lakes at the site were quite acidic and the lakes contained high concentrations of sulfate and dissolved metals. Surface runoff water from monitoring plots in 1976 were only moderately acidic and the mean concentrations of most dissolved metals and sulfate in the runoff waters were less than 25% of the concentrations found in Lake B. By 1980, vegetation had become well established over most of the mined area. However, the five final-cut lakes (Lakes A-E) remained acidic. The purpose of the study reported here was to determine the physical and geochemical subsurface conditions of the spoil materials, the hydrologic flow system within the spoils, and the magnitude of acidic spoil leachate migration into the lakes and into the wetland prairie north of the mine site. Lakes A, B, C, and E were still very acidic in 1980-1981 (pH range: 2.9-3.4; aridity range: 110-860 mg/L as CaCO3). Lake B had the worst water quality, with mean concentrations of dissolved aluminum, iron, manganese, and sulfate of 75, 25, 37, and 3340 mg/L, respectively. Shallow groundwaters in spoil materials around Lake B had a lower pH and about two to four times greater concentrations of aridity, aluminum, manganese, zinc, and other metals and nearly 20 times greater dissolved iron than the water in Lake B. The shallow groundwaters in undisturbed soils and bedrock 10-30 m north of the mine site spoils generally were alkaline, but did contain occasional moderately high sulfate and metal concentrations. Pumping tests and recovery tests conducted in piezometers showed hydraulic conductivities of spoil materials ranging from less than 10- to 10-5 m/s; hydraulic conductivities of natural soils and bedrock were measured between 10-10 to 10-7 m/s. Water levels in spoil materials increased rapidly in response to rainstorm events and the recession curves were relatively quick to decline, lasting only two or three days. Quick infiltration of rainfall into the spoil surface and percolation down to the water table caused a rise in groundwater levels and movement of spoil groundwater into the lakes. This groundwater component in the lake water budgets could account for most of the water quality problems occurring in the acid lakes.

Jan 1, 1984

By S. L. Bessinger

Introduction In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway&apos;s complex geometric cross section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Head losses in mine entries are calculated using Atkinson&apos;s Equation. [22H= KPLQ (English) H= KP 3Q (SO (1) 5.2AA] where H = pressure loss, in. of H2O (Pa); K = friction factor, lbf•min2/ft4 (kg/m3); P = perimeter, ft (m); L = airway length, ft (m); Q = airflow quantity, ft3/min (m3/sec); and A = flow cross-sectional area, ft2 (m2) In this equation, the friction factor, K, is an empirical constant that describes the aerodynamic roughness of the airway. Typically, the K-factor for a given airway is determined by measuring the factors H, P, L, Q and A in Equation (1) and calculating K. Tables of friction factors calculated in this way are found in textbooks and handbooks that deal with mine ventilation analysis. Unfortunately, very few K-factors have been measured on longwall faces, and the accuracies of those that have been measured are entirely site specific, because of the wide variety of equipment found on longwalls. The development of a technique for prediction without mine-site measurements of the friction factor for any longwall face, regardless of its configuration, will thus be very useful in the design of ventilation systems for mines in which longwall mining is practiced. Calculation of pressure losses using Atkinson&apos;s Equation (1) and empirically determined K-factors provides accurate and useful approximations in cases where the airways have relatively simple cross sections. However, a careful analysis using the principles of fluid mechanics shows that such calculations are based on two assumptions that are not strictly correct when there are obstructions in the airway. The first assumption is that the air velocity distribution in the cross section, particularly around the perimeter, is uniform. This assumption results from the fact that the tabulated K-factor values found in the literature are based on field measurements with uniform conditions. Such uniformity does not exist in longwall airflows. The second common assumption is that the K-factor, and corresponding head loss, is independent of the Reynolds Number (NR) for a given flow. In fact, this assumption is not strictly correct, and is particularly erroneous in the case of irregular protuberances into the airflow, such as those found on a longwall face. The errors arising from the assumptions may be avoided by using K-factors calculated by a newly devised method, described below, which takes into account the fundamental principles of aerodynamic drag analysis. This new technique has two advantages: first, it is flexible enough to model any longwall, regardless of equipment configuration; second, it employs terminology and equations familiar to those who perform mine ventilation analysis, using K-factors, for which ventilation engineers have an intuitive understanding, rather than drag coefficients. To provide guidance for development of a longwall drag model, data were taken on two modern longwalls operating in substantially different conditions. Pressure measurements at Mine B were made with 200-foot (61-m) sections of 1/8-in. (3-mm) diameter plastic tubing, attached to a Dwyer Magnehelic gauge. Pressure drops were measured in 200-foot increments down the face, and summed to give the drop for the entire face length. This technique was found to produce small, repeated errors because of the number of segments required to span the longwall. At Mine A this problem was avoided by using a single, continuous, plastic tube for the entire face length. The psychrometric properties of the air were measured for both Mines A and B. A calibrated, standard vane-anemometer was used to measure the airflow on both faces. Finally, numerous dimensions were measured on both faces, and face profile drawings were obtained to allow detailed evaluation of the face equipment geometry. From this information, accurate evaluations of the average wetted perimeter and average area of the longwall face airways were made. Since the airflow is not confined to inside the powered supports at all points along the face, a quadratically weighted average of the airflows measured at various stations along the face was calculated: [n2Qavg =Qi Ii / It(2)i=1] where [Q, avg = average airflow for analytical purposes, ft3/min (m3/sec); Q= airflow at station i, ft3/min (m3/sec); 1= length of segment represented by Q, ft (m); h= length of longwall face, ft (m): and n= number of quantity measurement stations.] The quadratic weighting scheme was chosen because the

Jan 1, 1992