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By Joseph E. Worthington

The announcement of a high-grade ore intercept near Wenatchee, Chelan County, WA, by Asamera Minerals in 1983 triggered widespread interest in this old gold mining district. The district has a long history of exploration and production and was actually one of the few underground gold mines that operated in the post World War I1 years in the US. Gold occurs in a north-trending structural zone in folded and faulted Eocene arkose intruded by andesite and dacite dikes having dates in the 42 to 50 m.y. range. Ore has been mined selectively from ladder veins perpendicular to the trend, from the south end of the trend near Squilchuck Creek. These veins narrow and silicification becomes pervasive northerly where wider zones were caved later in the operation. Recent discoveries at "B" reef at the north end are also in wide zones of silicified arkose. The gold lodes originally cropped out as prominent silicified and iron-stained ridges that were described in the early days as "reefs." The early discoveries were made in out- crop about a hundred years ago, and two claims, the Gold King and Macbeth lodes, covering the most prominent out- crops on the north side of Squilchuck Creek, were staked in 1885 and patented shortly thereafter. A five stamp mill was erected in 1894, but failed to liberate gold for amalgamation. The two patented claims were acquired in 1934 by J. Keegan, who carried out limited exploration and shipped small amounts of flux ore to Tacoma during the next few years. There was no substantial exploration or development of the claims, however, for several years after that time, and ownership of the claims eventually passed to a series of trusts administered by a bank in Seattle.

Jan 1, 1990

By Yoginder P. Chugh, M. Silverman

Premining investigations involve the determination of roof, floor, and coal seam characteristics to delineate areas of troublesome ground control prior to development of a mine. Data obtained can be useful in mine design, equipment selection, estimation of face productivity, development of ground control plans and assessing ground control costs, and in conducting feasibility studies and financial analyses. This paper discusses 1) the causative factors associated with ground instability, 2) how useful data can be obtained for these factors during premining investigations, and 3) analysis techniques to delineate probable areas of ground instability.

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Square-set timbering of stopes developed from stull support methods because veins too wide for stulling were being mined. In general, the void caused by the stoping of the ore is filled with a system of interlocking timbers. Timber framework supports the wall rock, the mining operations, and provides access. A timbered stoping method can be used in all widths of veins at any dip. The timbering can be normal and parallel to the planes of the walls or at angles to these planes. Irregularly shaped ore bodies can be worked by timbered methods. Pinching or swelling of the vein will not adversely affect the method. Timbered stoping is especially advanta¬geous where weak ground is encountered in the vein or the wall rock. The ground is unsupported only until timber sets are stood. These stopes can be filled with sand or waste when desirable. The customary method of developing a square-set timbered stope with haulage drifts above and below and timbered raises on each endline is shown in Fig. 1. The drifts are timbered the full stoping length. The raises have two compartments with the manway and timber slide in one compartment and the chute in the other. The drifts and the raises cannot be widely spaced as handling timbers is an important consideration. Stoping can begin immediately above the drift timber and continue in both directions and upward from the raise. Breasting down each blast round for only one timber set laterally is the common practice. The blasted ore falls to a lower floor where it is slushed or trammed to the chute. Drilling can proceed on the upper floor while broken ore is being removed from the stope. As shown in Fig. 2, square-set timbering may be used in cut-and-fill stopes to support the back and walls until fill is added. There are various ways of developing an ore block for timbered stoping. Fig. 3 depicts the situation for the main haulage in the vein with an intermediate drift separated from it by a sill pillar. The ore is slushed into the chute connecting the two drifts. The situation in Fig. 4 is identical to that in Fig. 3 except that no sill pillar is used. In Fig. 5, the haulage drift development is done in the footwall with crosscuts driven to intersect the ore body. The intermediate drift development is above the crosscuts with a minimal sill pillar. The main bottleneck in timbered stoping is in the timbering process. Drilling in the stope is generally held up while timbering is done. Two or three men are re¬quired to do the timbering and an additional one or two men are needed to transport and raise or lower the tim¬ber. Good organization is essential to timbering. Skill must be exercised in standing the timber to support the ground and still maintain good production. COSTS AND PRODUCTIVITIES Development and stoping costs are presented in Tables 1 through 7 for the situations in Figs. 3 and 5,

Jan 1, 1982

By A. Wahab Khair, Negendra P. Reddy

INTRODUCTION The continuous mining machines which were introduced in the 1950's now account for more than half the production of coal from underground mines. Unfortunately, these continuous miners which were designed for increased productivity have also increased the concentration of respirable dust in mines. Inhalation of this respirable dust by miners causes the Coal Worker's Pneumoconiosis (CWP). Concern for CWP increased in the United States following a 1963-65 survey that showed definite x-ray evidence of CWP in 9.8 percent of working miners and in 18.2 percent of nonworking miners in the Appalachian bituminous coal region (Breslin and Niewiadomski, 1982). One of the major objectives of the Federal Coal Mine Health and Safety Act of 1969 was the prevention of CUP. This law initially limited the daily average exposure of coal miners to 3 mg/m3 of respirable dust. However, following the results of a probabilistic model developed by Jacobsen, et. al. in 1971, the limit was reduced to a mass concentration of 2 mg/m3 unless quartz content exceeds 5 percent, a rare occurrence in mines. Based on x-ray categorization prior to death it is believed that the mass concentration in the respirable fraction is a parameter of primary interest for hygenic evaluation of dust exposure. Elorgan, Burgess and Jacobson (1973) suggested that the quantity of inhaled dust, although influential is not the only aspect of exposure which is significant in producing CWP. Miners working in higher rank coal (anthracite) mines are more susceptible to CWP (Morgan, Burgess and Jacobson, 1973). This also confirms the findings of the British (Bennett, Dick and Kaplan, 1979 and Hurley, et. al., 1979) and Germany (Reisner and Rubock, 1977). Based on the cy to toxicological evidence Reisner and Rubock (1979) suggested that when coal properties are similar, increasing quartz and mineral content increases toxicity. David, Ottery and Le ?oux (1975) examined that as the severity of the CWP increased the retention of the coal dust content in the pulmonary region remained the same, while that of quartz content increased significantly . Leiteritz, Baur and Bruckman (1971) studied the various components of coal dust and showed that coal particles comprised a smaller fraction of the mass of fine dust than did particles , of quartz and minerals. It is ironic that quartz and mineral matter both comprise smaller fractions of the higher rank coal dust (Dodgson, et. al., 1971 and Dwiggins, 1981). Results of a recent research revealed that quartz or any other mineral examined had no strong or consistent effect on the incidence or progress of CWP (Attfield,

Jan 1, 1986

By Frederick K. A1lgaier

The Bureau of Mines' Denver Research Center initiated a subsidence research program in 1978. The short ten goals of the program were to determine practical methods and equipment for measuring subsidence and demonstrate these methods by monitoring subsidence at several mines having mining, geologic and topographic variables typical of those encountered in the western United States. Data from the monitoring program would be supplied to cooperating mining companies and used by the Bureau to accomplish the long term objective of developing and validating subsidence prediction techniques applicable to mines in the West. Subsidence monitoring is being conducted at five mines in Colorado and Utah with substantially different mine plans, mining methods, depths of cover, and topography. The monitoring procedures used at each site are similar and include designing the monitoring network layout, installing the subsidence monuments, initial and periodic surveying of the monuments, and processing the survey data. The network layout, surveying methods, and types of monuments vary according to the conditions at the individual sites and the type of data being collected.

Jan 1, 1982

By B. Salopek

The greatest part of crushed stone, gravel and sand production is extracted through open pit operations, whereas gravel and sand are also partially extracted from riverbeds. Its obvious that such a mining activity provokes some changes in the configuration and environmental characteristics of the region, in the temporary or permament disapperance of vegetable or animal species and in the creation of noise, dust and waste water. However, the above-mentioned mining operations shouldn not have only negative effects on the landscape and land potential, but they can discover and predetermine the potential of the exploited areas for various communal purposes after exploitation (eg. sports and recreation centers, campsites, resorts, fish-farms, etc.). In this paper, the production of construction aggregates in plants located in the vicinity of Zagreb, the capital of Croatia is presented. The advantages and disadvantages of aggregates? production in urban areas in the spirit of sustainable development are discussed.

Jan 1, 2003

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Shrinkage stope mining can be employed in steeply dipping veins where both the wall and the vein material are sufficiently strong to stand with only minimal ground support. Shrinkage stope mining takes advantage of the swell in volume produced in breaking in-place ore. This expanded volume of ore is allowed to occupy part of the mined-out vein, and in so doing provides ground sup¬port and a working surface for the miners. Only enough of the ore is drawn after each stoping cut to allow work¬ing room. Not until the entire mining block has been broken is the full amount of broken ore drawn from the stope. DEVELOPMENT PRACTICES FOR SHRINKAGE STOPING The vein is developed into stoping blocks, each of which is 76 to 91 m (250 to 300 ft) in length and 61 m (200 ft) between levels. Level development, described later, is either on the vein or parallel to it with crosscuts periodically driven over to the vein. There are usually two exits from each stope. The first entry is a raise at one end driven either conventionally by raise climber, or by a raised borehole. Present practice is that this raise, put in from level to level, is either a 2.1- x 2.1-m (7- x 7-ft) Alimak raise or a 1.2- to 1.8-m (4- to 6-ft) diam borehole. Either raise system will require pro¬visions for an ore chute carried with the stope to the haulage level or the scram drift. At 9- to 12-m (30- to 40-ft) vertical intervals, dogholes are developed from the raise so that when stoping breaks back into these dogholes, timber or steel sets and necessary service facilities can be installed. The second entry to the stope may be either a similar raise at the stope endline, or more economically, a cribbed raise carried with the stope. As previously stated, development of the ore block is achieved by either drifting on the vein or by paralleling the vein with a footwall lateral and then crosscutting over to the vein at specified intervals. In the following discussion, three basic cases are illustrated in Figs. 1, 2, and 3. In Fig. 1, development is in the footwall by a 3- x 3-m (10- x 10-ft) trackless lateral. A heading is also driven on the ore and taken either full width or driven a standard size and slabbed to the vein width. Approxi¬mately 9-m (30-ft) long crosscuts on 7.6-m (25-ft) cen¬ters are driven from the lateral to the vein. For ease in muck handling they are usually put in on a slant rather than at right angles. A single orepass, located off the lateral and in the center of one stope block, can serve not only that stope, but also one on either side, up to a tramming distance of not more than 137 m (450 ft). Raises are put at the stope endlines on 91-m (300-ft) centers. The back of the vein is shot down and just enough broken material is drawn off to permit access for men to work off the broken ore. Development in Fig. 2 progresses by driving a 3- x 3-m (10- x 10-ft) trackless heading on the footwall side of the vein. The ore block is delineated by raises on 91-m (300-ft) centers. Crosscuts are again driven on 7.6-m (25-ft) centers over to the vein. Approximately four rounds are then taken up to the vein and each round is successively belled out so that all the crosscuts (pockets, in this case) connect. In this development, as opposed to that in Fig. 1 where there were no belled

Jan 1, 1982

By M. J. Lowe

National Institute for Occupational Safety and Health (NIOSH) studies show that 43.5% of surveyed coal preparation plant workers had noise exposures exceeding the Mine Safety and Health Administration Permissible Exposure Level. Sound levels around vibrating screens in these plants often exceed 90 dB(A). NIOSH is currently developing noise controls for horizontal vibrating screens. To characterize noise sources, NIOSH researchers performed sound pressure level (SPL) measurements on a vibrating screen at their Pittsburgh Research Laboratory. The results show that the entire screen contributes to noise below 1 kHz and the vibration mechanism housings are most significant above 1 kHz. Constrained layer damping (CLD) treatments and an enclosure were used to reduce mechanism housing noise. These were evaluated using sound power level measurements according to ISO 3744. The CLD treatments reduced the A-weighted sound power level by 3.1 dB in the 1 to 10 kHz one-third-octave bands. A panel-on-frame vibration mechanism enclosure using various types of panels further reduced the A-weighted sound power level from the CLD configuration in the 1 to 10 kHz one-third-octave bands by 3.7, 4.0, and 3.9 dB for aluminum, steel, and Dynalam panels, respectively. The combination yielded a 7 dB reduction from baseline in A-weighted sound power for the same frequency range.

Jan 1, 2010

By J. Herbst, A. J. Lynch, E. V. Manlapig, G. C. Throne, J. S. Hall, G. Poling

Modelling of industrial sulphide flotation based on data collected in several flotation plants is discussed. A mechanistic approach is used in which it is recognised that flotation is a two stage process with concentration occurring both in the pulp and in the froth. The important features of the model, which is a kinetic model, are – 1. for each particle species there is a range of flotation rates which corresponds to the range of particle sizes, 2. the flotation rates of the liberated gangue particles which have no hydrophobic properties are dependent on the flotation rate of water and are controlled by variables such as the air and frother addition rates, and the pulp level, 3. as an approximation the valuable sulphide particles may be regarded as being present in two groups with high and low flotation rate constants, 4. increase in the addition rate of collector tends to increase the proportion of sulphide particles which are in the high flotation rate constant group. The actual flotation rate of each group of particles is a combination of an "inherent" flotation rate and a flotation rate which is dependent on the flotation rate of water, at particular air and frother addition rates and pulp levels there is a physical limit to the rate at which particles can be transported into the concentrate and this is recognised in the

Jan 1, 1976

By Ralph H. Knode, Stephen P. Collings

INTRODUCTION The Crow Butte uranium deposit is located in northwest Nebraska near the town of Crawford (Figure 1). The deposit was discovered during the fall of 1980 by Wyoming Fuel Company, operator of the Crow Butte Joint Venture. The Crow Butte Joint Venture is owned 50 percent by Wyoming Fuel Company, a subsidiary of KN Energy, Inc.; formerly Kansas-Nebraska Natural Gas Company, Inc., 40 percent by Ferret Exploration Company, Inc., and 10 percent by First Exploration Company. A preliminary announcement made in January 1981 indicated a "probable potential" reserve in excess of 25,000, 000 pounds U303. Drilling during 1981 and 1982 confirmed these reserves and indicated the presence of more than 30,000,000 pounds U308 having an average grade in excess of 0.25 percent U308. Wyoming Fuel Company was initially attracted to the area in 1978 by favorable regional geology for sandstone uranium deposits and weak radioactivity noted in regional oil and gas holes. Exploration drilling began in 1979 following the formation of the Crow 3utte Joint Venture. Drilling has continued up to the present, at which time the Joint Venture is ready to proceed with a pilot scale solution mine. Environmental baseline data have been acquired and the permit applications for the pilot scale mine have been submitted to the Nuclear Regulatory Commission and to the applicable state agencies in Nebraska.

Jan 1, 1984

By M. K. Gangal, E. D. Dainty

The paper describes the use of portable aerosol monitors in determining real-time concentrations of diesel soot and respirable total dust in dieselized mines. The field studies were conducted in a simulated diesel environment in the laboratory where no other type of dust was present, and in four underground mines with multi-diesel machine operations. Two packages for the real-time monitoring of diesel soot were prepared which included two types of aerosol monitors: a Miniram, model PDM-3, and a Sibata, model PDS-1. These monitors utilize the principle of light-scattering to measure the airborne mass concentration in the environment. The monitoring device results were compared against the Time-Weighted-Average (TWA) values produced by the gravimetric filter sampling method, permitting calibration factors to be determined for each field test. Both instrument packages provided comparable results. The instrument calibration factors varied for different tests, which may be due to different material composition, particle size distribution, density and shape of the aerosols. This approach for real-time determination of diesel soot is found to be simple, inexpensive and easy to use. More field tests for diesel soot determination and further research on size distribution of particulates in dieselized mines are required.

Jan 1, 1989

By L. A. Puchkov

The paper presents the outline of the results in the algorithms and software of the microprocessor-based system for the powerful drugline bucket automatic motion, as well as the technological process control and functional diagnostics of the electrical equipment. The system is, aimed at the increasing both the draglines' working efficiency and the professional, competence of the operators. The excavator-dragline, equipped by the automatized operation, diagnostic and information processing system, possesses considerable technological: flexibility, which can widen therefore its field of application. The tasks of the dragline bucket motion optimal operation, programming of the excavation and overburden dumping processes, as well as the tasks of the control and the analysis of the technological parameters and the diagnostics of electrical equipment conditions, are defined in the paper. In our paper we present the results of these tasks being solved, providing the algorithms of the main mechanisms for the automatic operation for the swinging bucket controlled motion, the algorithms of the statistical analysis of the working efficiency parameters the algorithms of the diagnostics for electrical driving systems. The algorithms of the motion operation are constructed using the predicting dynamic models. The diagnostic algorithms at the normal working conditions form a number of algebraic invariants always equal to Ø with the accuracy up to when the electrical equipment is parametrically adequate. The algorithms of evaluating and analysis of the excavation efficiency make use of the direct measurement and analytical identification. The working condition and quality of the operation and control algorithms are investigated by simulation and natural testing. The results of the mathematical simulation of the bucket motion (with the 20 m3 capacity) of the excavator-dragline with the boom of 90 m in length are given in the paper.

Jan 1, 1992

By N. B. Aughenbaugh

A study was made to determine those petrographic parameters that are correlative with carbonate-aggregate quality as indicated by the Los Angeles abrasion and absorption tests. Samples collected at 28 sites in northern and eastern Indiana were analyzed petrographically, chiefly for void content, grain size, grain-size variation, grain shape, and grain interlock. Most of the sampled sites are in the belt of Silurian rocks of east-central and north- western Indiana. Though a few sites lie outside this belt, most of these have only a thin Devonian section above the Silurian. Three sites are non-Silurian and are included for comparison. Most of the samples studied are from Middle and Upper Silurian units. Both biohermal and nonbiohermal facies are included. The biohermal rocks are typically dense but vugular, finely crystalline granular dolomite. Nonbiohermal facies are porous, medium to coarsely crystalline, granular dolomite with dolomite rhombs The Los Angeles abrasion losses ranged from 20 to 46 per cent, the rocks in the eastern part of the state showing the higher values. Absorption tests indicate the same general quality trend. Comparison of petrographic studies with Los Angeles abrasion- and absorption-test results indicate the following correlations. (1) As per, cent of voids increases, quality decreases. (2) As average grain size increases, quality decreases. (3) Variations from mean grain size appear to have no effect on the quality. (4) As angularity and roughness of the grains increase, quality increases. (5) As grain-to-grain interlock increases, quality increases. Variations in the calcium-magnesium ratio have no apparent effect on aggregate quality, except as they affect the texture.

Jan 1, 1966

By Peikun Liu, Guangyu Sun, Raja V. Ramani, Shambhu Sharan, Raman Srikanth, Bharath K. Belle

The contribution of the various outby sources to the contamination of the panel intake air has been the subject of a relatively small number of studies. In this paper, after a brief review of past studies, results from recent experiments on rates of contamination of intake air by outby dust sources are presented. Present study indicates that each outby source by itself may not be a major contributor. The loading point in the continuous miner section did not appear to be of major concern from contamination point of view. Belt entry including transfer point contributes significantly to the intake air dust level, though reentrainment was not a major source of dust. In longwall panels, the stageloader-crusher was the most dominant source. The panel-to-main belt transfer point was not a major source. Taken collectively, the contribution of all outby sources can be significant. In one longwall section, at the beginning of the longwall face, the contribution of all the outby sources and the headgate cut-out operation was observed to be about 80% of the permissible exposure limit (2 mg/m3). In modem high production longwalls, the application of effective dust control measures to reduce the dust generated from the outby sources is essential to achieve compliance with the dust standards.

Jan 1, 1997

By Michel Grenier, Mahendra Gangal

Two methods used to measure the exposure of mine workers to diesel exhaust particulate were compared. One method was developed by the U.S. Bureau of Mines and the University of Minnesota's Particle Technology Laboratory. The method which was originally developed for use in coal mines is a size selective sampling technique. The other method uses an “ashing" approach to assess the exposure to diesel particulate. This second method is based on the principles of sampling according to the ACGIH respirable dust sampling criteria and the use of a silver membrane filter as collection substrate. Because a silver membrane filter is used, both respirable combustible dust (diesel particulate) and respirable silica can be measured from one sample. The comparison study took place in a Canadian lead-zinc mine during tests which were primarily designed to evaluate the gaseous and particulate emissions characteristics of new diesel engine technology. The study has shown that under the test conditions encountered, some differences were observed between the results as obtained by the two methods. These differences seem to depend on the characteristics of the respirable dust found in the work- place, i.e., size and composition. The impact, on sampling and analysis of diesel particulate, of the proposed guideline (ACGIH) for a TLV® of 0.15 mq/m3 is discussed briefly.

Jan 1, 1997

By A. D. S. Gillies

The effectiveness of coal mine face ventilation and the ability of air flow to maintain satisfactory methane and respirable dust levels in the working face area has been studied through the use of scale modeling techniques and underground testing procedures. In particular, consideration has been given to examination of the effects of operation of a high air flow machine mounted dust scrubber unit on face air patterns. Results obtained through a cooperative study with the Peabody Coal Company demonstrate that new mining cycles can be considered which increase production efficiency while maintaining satisfactory ventilation conditions.

Jan 1, 1982

By James P. Connell

HISTORICAL BACKGROUND The American Concrete Institute defines shotcrete as "mortar or concrete conveyed through a hose and pneumatically projected at high velocity onto a surface." This definition thus includes what is traditionally known as gunite, which is a pneumatically applied mortar. In mining practice, the term shotcrete is restricted to pneumatically applied concrete, and this differentiation will be used in this chapter. In 1914, following the invention of the mortar gun in 1907, then chief engineer of the US Bureau of Mines (USBM) George Rice developed the gunite process for underground test work at the USBM facility at Bruceton, PA. After World War I, gunite was used extensively in American mines and was also utilized for underground civil works such as the San Jacinto tunnel in California. The greatest development was in Europe where, as early as 1911, gunite was successfully used as an overlay for deteriorated tunnel linings. In 1951, the Swiss firm Aliva developed a pneumatic gun capable of handling coarse aggregate, thus making possible the first use of shotcrete at the Maggia hydropower development. Initially, shotcrete was used to reduce manpower requirements for forming and placing conventional concrete. However, by 1954 Sonderegger was reporting that the structural advantages of shotcrete were derived from its flexibility and from the fact that it could be applied almost immediately after the opening had been made. The incorporation of wire mesh into the shotcrete led to the new Austrian tunnel method or NATM. The use of shotcrete in American mines has been implemented more recently. This delay seems to be due to previously unsuccessful experiences with gunite as a structural material and to the US reliance on wood or steel supports in main-line haulageways. The long experience with the apparently more substantial rigid supports led mine operators to be reluctant to accept the new and seemingly unrealistic lighter shotcrete support. APPLICATION REQUIREMENTS Shotcrete is a relatively new material for use in underground support systems. Consequently, experienced miners are not always available who are capable of applying the material effectively. Shotcrete, particularly in the small cross sections typical of mine shafts or haulageways, is applied in cramped quarters under less than ideal conditions. Adequate lighting should be made available. The surface should be clean and free of running or dripping water. It may be necessary to collect flowing water in plastic pipes or water collection devices. Any dry cement dust from previous shotcrete applications should be washed from the surface in order to assure a good bond. The US Bureau of Reclamation (USBR) while shooting test panels at the Cunningham tunnel in 1974, found that experienced shotcrete operators were able to obtain up to three times greater compressive strengths than were obtained by unskilled operators using the same equipment and shotcrete mix. ENVIRONMENTAL AND SAFETY REQUIREMENTS Since sodium and potassium hydroxide, as well as other moderately toxic compounds, are often contained in shotcrete (particularly where accelerators are used), safety precautions must be taken to prevent skin and respiratory irritation. Nozzlemen and helpers are required to wear gloves, protective clothing, and ventilation hoods with a filtered air supply. Respirators approved by USBM, equipped with chemical filters that will not pass the caustic mists, may be permitted in lieu of hoods if goggles or safety glasses are worn. Protective ointments are available to reduce skin irritation. All air and shotcrete feed hoses should be equipped with safety-type couplings and secured with safety chains at each coupling to prevent whipping in the event of a hose or coupling failure. Some environmental effects can take place down-stream from the development face being supported. The accelerator compounds, as well as the portland cement used in the shotcrete, will be found in the rebound material which falls to the invert of the heading. Since these compounds may be leached from the rebound material and carried by the drainage system, it may be necessary to install neutralizing or other water treatment facilities. Investigations may find that the final reaction with other compounds being leached from the mining operations may result in a more or less environmentally acceptable end product. USES OF SHOTCRETE General Uses Shotcrete, as a combination of cement, aggregate, and accelerator, is utilized for underground openings such as shafts, adits, haulageways, and service chambers for the following general purposes : (1) primary sup¬port; (2) final lining; (3) protective covering for excavated surfaces that are altered when exposed to air (the protective covering may be of a temporary or final nature); (4) protective covering for steel or wooden supports, rockbolts and rockbolt plates, heads, nuts, and other mats, including wire fabric, used to prevent rock-falls; and (5) as a lagging material in place of timber, steel, or concrete between steel or wooden supports. These applications can be grouped into three general use categories: shotcrete used as a rock sealant, shotcrete used as a safety measure, and shotcrete used as a structural support. Use as a Rock Sealant Thin applications of shotcrete can reduce or prevent slaking of shales or other rocks that are altered when exposed to the wetting and drying cycles created by mine ventilation circuits. While shotcrete may be effective in preventing such rock alteration, at the present time it is not as economical or efficient as other commercial sealants. However, if the sealant property can be incorporated into the structural support capability, the added contribution is usually helpful. Thin applications are not usually sufficient if the alteration of the

Jan 1, 1982

By D. T. McMurray

INTRODUCTION Shabanie mine, situated some 180 km east of Bula¬wayo, has been a producer of chrysotile asbestos for more than 50 years. The ore bodies occur in serpentinized dunite, which overlies talc-carbonate schist. A zone of relatively competent rock of varying thickness occurs between the schist and the ore bodies, which are gen¬erally less competent. The hanging wall of the ore bodies is economic, and the hanging-wall serpentine carries a variable subeconomic amount of fiber. It is important to note that, in general, the ore body competence is less than that of the foot and hanging-wall formations. Historical After surface operations ceased, cut-and-fill stoping was used to win ore from underground; this was success¬ful until the increasingly stoped-out area caused insta¬bility in the stope pillars and back. Consequently, dur¬ing the early 1950s, a gradual change to cave-mining methods was made, the ore being won by hand lashing in drawpoints, situated in the basement of the stope blocks, and passed through orepasses under gravity to the haulage level some 13 m below. About this time, interest was focused on the sub¬level caving method in use in Swedish iron ore mines: it was felt that it might be applied economically to the Shabanie ore bodies. Accordingly, in 1958, an experi¬mental stope block was laid out in which sublevel inter¬vals and extraction tunnel spacing were 9 m. The tun¬nels (ring drilling drives) were oriented on strike-in contrast to the Swedish system, in which crosscuts that retreat from hanging to footwall are used. The advantages of the method were quickly appre¬ciated by the operating personnel and, despite the in¬evitable teething troubles pertaining to the introduction of any new mining method, it was not long before sub¬level caving was providing a high proportion of the mill feed. The disadvantages also became apparent at an early stage, however, and, from that time to the present, continuing modifications have been made to mining lay¬outs in an effort to improve ore recovery. GENERAL DESCRIPTION OF METHOD The mine is served by a vertical hoisting shaft, in which two skips, a man cage and a service cage, provide adequate capacity for production requirements. The rock hoist is a Ward Leonard control hoist, in which two electric motors drive a common gearbox. The man winder is driven by an a-c motor. Several auxiliary shafts provide secondary egress and intake and return ventilation. Main haulage levels are above (Fig. I a and b). Blasthole fan patterns are drilled by drifters of 100 mm bore, drilling 41-mm holes; when a sufficient strike length has been drilled, a slot is cut in the upper¬most sublevel and the rings are broken into the slot. Initially, a limited tonnage is drawn, since it is essential to ensure that the hanging wall caves behind the retreat¬ing stope face. Once this has been established, maxi¬mum tonnage can be drawn, as described later in this chapter, under the heading "Draw Control." The broken rock is loaded by 0.14 and 0.20-m3 load¬ers into cocopans (rocker-dumping type of tipping truck), which are hand trammed to orepasses, discharg¬ing on the haulage where 11-t electric trolley locomo¬tives haul 3.95-m3 Granby cars to the main shaft bins. As is evident from Fig. 1, the layout is simple, the block is brought rapidly into production, there is a high degree of selectivity and flexibility, and the result is a low-cost high-productivity mining method. DEVELOPMENT Main haulages are developed at 3.2 x 3.2 m, and once the service winze connections have been completed the development of the sublevels is undertaken. The footwall drives are cut first, to obtain access to the block. These ends are of the standard section, 2.4 x 2.8 m, and from them crosscuts at intervals of 70 m are driven through the ore body to the hanging wall. These crosscuts are used to supplement the geo¬logical information previously obtained from diamond core drilling, and they provide additional and more de¬tailed data on fiber percentages and lengths, structural features, and other relevant criteria which are used to build up the geological assessment of the area and to classify it in terms of the geomechanics rock classification (Laubscher and Taylor, 1977). The crosscuts also allow the necessary orepasses to be sited conveniently so that tramming distances from the loading points are not excessive. Development Drilling Once the skeleton development has been completed, the extraction headings are developed at 2.4 x 2.4 m as shown in Fig. 1. Standard development practice is to use crews of a machine operator and his helper, equipped with air-leg mounted jackhammers, to drill rounds of 1.8 m with integral tungsten carbide tipped drill steel. The round drilled is a normal drag round, as shown in Fig. 2, but considerable attention is paid to the drilling of the perimeter holes to use effectively the

Jan 1, 1982

By S. Ersayin, I. Iwasaki

The application of a magnetic field was shown to be effective in controlling iron losses in a pilot-scale cationic silica fIotation in the processing of magnetic taconite concentrates. The design of a magnetic field distribution device and batch flotation test results using a 1.42-m3 (50 cu ft) WEMCO flotation cell are reported. Major losses of iron units in froth products were in the -25-µm (-500-mesh) fraction, and the application of a magnetic field decreased the flotation of fine magnetite, thereby improving the selectivity of separation. The "as-received" sample was sufficiently magnetized, and no further benefit was gained by the magnetizing treatment. The selectivity was somewhat adversely affected by demagnetizing the sample, though its effect was minimal due to the magnetic field. The device is simple in construction, low cost and may be readily installed in existing equipment. Key words: Iron ore, Flotation, Taconite concentrates, Cationic silica flotation, Magnetic field, Flexible magnetic sheet, Magnetizing, Demagnetizing

Jan 1, 2003

By L. J. Lindsay

Kaiser Resources Ltd. operates a coal mining complex in Southeastern British Columbia. A large surface mine, a conventional underground mine, and an innovative hydraulic underground mine are presently in operation. The products of these mines are fed to the Elkview plant. This coal has to be prepared to 9. 5% ash and 6'o total moisture for export from the port of Roberts Bank near Vancouver, British Columbia to Japan at a rate of 4. 5 Mltpy. The moisture requirement for the coal product was recently increased to 8. 0% total moisture. The entire operation has been described in various publications (1), (2), (3). Details of the venture will not be repeated in this paper. In this paper we hope to describe the modifications made to the original coal preparation plant in order to satisfy market requirements for quality and quantity of coal.

Jan 1, 1974