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By E. Scott Bair

Renewed interest in the vast reserves of anthracite coal in eastern Pennsylvania led to the use of a numerical model for simulating the hydrologic effects of constructing an open-pit anthracite mine. The proposed mine would be approximately 1800-m (5900-ft) long, 1000-m (3300-ft) wide, and 300-m (1000-ft) deep. A two-dimensional, finite-difference, ground-water flow model was used to calculate ground-water seepage rates into the proposed mine, to evaluate the effectiveness of various deep-well dewatering systems, to calculate drawdown in adjacent wells and properties, and to estimate the capacity of the water-treatment plant. Site-specific values of pertinent hydrogeologic parameters were used in the simulations.

Jan 1, 1980

By Michael L. Clar

Reclamation of surface mined lands has two basic objectives: (1) the elimination or abatement of undesirable environmental conditions, and (2) the restoration of mined land to its former or more productive uses. This paper addresses some of the planning concepts and site factors that influence the selection of suitable land uses for surface mined lands. Three planning concepts included in the discussion are: (1) mine land planning, (2) comprehensive land use planning, and (3) environmental site planning. The site factors that will influence the selection of an appropriate land use have been grouped into three categories: (1) legal, (2) geographic, and (3) natural. The natural factors are further subdivided into three groupings: geomorphology, soils, and climate.

Jan 1, 1981

By C. C. Graham

In this paper, the problems which occurred on introduction of dense medium cyclone coal washing in Australia are reviewed together with the subsequent measures taken to obtain a satisfactory magnetite. Specification and testing of magnetite for exploration, quality control and in-plant check purposes is discussed. The disposition of Australian magnetite deposits in relation to the main coal producing areas is examined, and comments are made on the economics of mining, milling and beneficiating magnetite for coal preparation uses alone. Future trends in the Australian Coal Industry are predicted, and the need for a planned approach for the future is demons traced.

Jan 1, 1982

By Ottmar Burghardt

Sintering of iron ore has undergone a dramatic development world-wide. Today, the annual world production amounts to approximately 600 million tons. At present, the sintering process in competition with pelletizing, seems to offer certain advantages inasmuch as it allows the use of solid fuels. Also, its charging behaviour in the blast furnace is, in general, considered as being better than that of pellets. This is true, in particular, for the charging behaviour and the gas permeability in relation to the grain size. Moreover, instead of charging limestone into the blast furnace, for instance, the basic components which are necessary for the burden materials can generally be supplied in a simpler and more cost-effective way by using basic sinter instead of basic pellets. While in sintering, with basicity rising to the range of 1.6 to 1.8 CaO : SiO2, the capacity increases and the sinter quality shows improvements in most of its characteristics, the pro¬duction of basic pellets generally requires a higher fuel rate, when compared to acid pellets, while production capacity decreases and some of the quality characteristics deteriorate.

Jan 1, 1982

By A. W. Khair

This paper presents an analysis of respirabie dust generation due to the action of a continuous miner. Underground coal cutting by a drum-type continuous miner was simulated in the laboratory using a specially designed unique automated rotary coal cutting simulator (ARCCS). An analysis of -400 mesh particles was done on gravity collected material. Samples of entrained dust were collected through the use of cascade impactors. Both samples were analyzed for size distribution and particle shape. Using the simulated coal cutter permitted variation of operating and in-situ parameters to provide information on the effects on generation of -400 mesh particles. In these tests coal blocks with an approximate dimension of 18 in. x 15 in. x 6 in. (45.7 cm x 38.1 cm x 15.2 cm) were first subjected to confining pressures, equivalent to in-situ conditions. Such blocks were then cut by the cutting head of the ARCCS, thus simulating the action of the continuous miner in underground coal mining. During the tests under a particular set of in-situ and operating parameters a number of other parameters such as 1) penetration of bit into coal, 2) penetration resistance (thrust and cutting pressures), 3) rotating velocity of cutting head, and 4) acoustic emission activity in the coal block were monitored. After each cutting cycle the fractured surfaces were photographed and a velocity survey was conducted by using a sonic technique. At the end of each experiment the cutting paths of the bits in coal were photographed using an optical microscope with an attached camera.

Jan 1, 1985

By Tan Zhenxiang

INTRODUCTION Koolyanobbing is an open cut iron ore mine, which is located in Yilgarn region of Western Australia. To prevent the operators from suffering silicosis diseases and protect the air from being polluted, a series of dust control measures have been taken. These measures include ventilation and dust control system for sample preparation shed, dust collecting system for drill rig, dust suppression system for ore processing plant, dust control for haul road. The details related to individual dust control measure are dealt as follows. SAMPLE PREPARATION SHED DUST CONTROL The sample preparation is one of the major dust sources at Koolyanobbing. The activity is being carried out at the sample preparation shed. It typically needs to go through the following stages, that is, crushing, splitting, drying and pulverising. The amount of dust generated from crushing and splitting is mainly determined by the moisture content contained in the pre- pared sample. A significant amount of dust is generated at the sample pulverising stage due to cleaning sample bowl with com- pressed air. To protect the sample preparing personnel's health and make the dust level in the air meet the requirement from the health and safety authority, a central ventilation and dust control system has been installed at the sample preparation shed. As shown in Fig. l, the system consists of suction heads, suction lines, dust collector and vacuum fan. The fan creates vacuum environment in the system, therefore the dust is drawn into the dust collector through the suction heads and suction lines. The dust collector is of fabric filter. When the dusty air comes to the filter, the dust in the air is collected and retained on the surface of the filter while the clean air passes through and enters the atmosphere. The retained dust is blowed out by the pressured air pulses and deposits into the dust bin underneath the filter. A time relay device is used to control the cleaning time and the cleaning interval. DRILL RIG DUST CONTROL A hydraulic percussion drill is used to drill blast holes. To carry out the drilling operation properly and efficiently, the cut- tings are being flushed out of the hole by the compressed air. In order to prevent from the fine particles in the cuttings escaping to the air, a dust collector is attached to the drill. The collector mainly consists of suction head, cyclone and bag filter. Refer to Fig. 2. The suction head is installed at the lower end of the feed beams and is movable. The moving range is up to 250 mm. It is not only used to restrict the cuttings, but also used to prevent water from entering the filters when drilling wet hole. A 406 mm diameter primary cyclone is used to separate the heaviest particles in the cuttings. It is also used as blast holes sampling equipment. One part of the dust particles are collected by a sample bag through a splitter attached to the end of the cy- clone. The rest of the particles fall down to the ground through the closing rubber. The bag filter is driven by a hydraulic motor (1). It is used to collect the lighter particles in the cuttings. The filtering area is 16 m2 and the dust separation efficiency is up to 99.9%. The working principle of the filter is demonstrated in Fig. 3. When the rock drill (9) flushing is turned on, the pressurised air from the compressor is directed to the rock drill and to the pres- sure switch (10). The pressured air will open the pressure switch to de-energise and close the solenoid valve (4) The pressure affecting the cut-off cylinder (5) keeps the exhaust port flap open. When the flushing valve (8) is closed, the pressure affecting the cut-off cylinder disappears and the internal intake opening flap is closed by a spring. The pressure also ceases to affect the pressure switch and the spring switches the contact to the point 2. The time relay (11) is now activated and gives a 0.5 second current impulse to the solenoid valve and keeps the current cut off for approx 8 seconds. When the solenoid valve is live, it opens and lets an impulse of pressured air to the filters (2). These impulses will clean the filters.

Jan 1, 1997

By R. E. Le Messurier

GENERAL DESCRIPTION The mines are located at Broken Hill in the western part of New South Wales, Australia, 930 km west¬northwest of Sydney. Mining Methods Where pillars must be left between stopes, extraction of a block of ore between two or more levels is carried out in stages: 1) Stopes are mined overhand by horizontal cut¬and-fill stoping, or by long-hole stoping. The various methods of horizontal cut-and-fill stoping used are desig¬nated as: -Loader cut-and-fill stoping with horizontal blast¬holes and diesel load-haul-dump (LHD) units. -Open stoping with horizontal blastholes and com¬pressed air or electric scrapers, or with diesel or electric LHD units. Square-set timber stoping, where ground condition precludes open stoping. The backs are sup¬ported with timber square sets. Pattern rockbolting of the backs may be used as an intermediate stage between open stoping and square-set stoping. 2) Vertical pillars are mined by undercut-and-fill stoping or by vertical crater retreat stoping. 3) Where a horizontal pillar has been formed below the level above the primary stope, this level pillar, usually 6.1 to 8.2 m in depth, may be mined by over¬hand square-set stoping or by undercut-and-fill. Long-hole stoping is used at the Zinc Corp. (ZC) and New Broken Hill Consolidated (NBHC) to mine parts of the zinc-rich A and B lodes. Drilling is carried out from drill drives and, in some cases, sills. Hole sizes used are 165 mm (61/2 in.), 110 mm (4'h in.), and 65 mm (21/2 in.) nominal diam with hole lengths being limited to 50, 30, and 20 m respectively. Ammonium nitrate-fuel oil (ANFO) and water gel explosives are used. Broken ore is loaded from drawpoints at the bottom of the stope using rubber-tired diesel-driven LHD units which deliver ore to an orepass. Annual production at Zinc Corp. and NBHC accord¬ing to mining method is given in Table 1. At Zinc Corp. mining is being carried out at depths varying between 200 and 832 m. At NBHC it ranges from 528 to 915 m below the surface. The average travel time to the stope is 30 to 45 min. GENERAL ORE BODY REQUIREMENTS Size, Shape, and Dip There are six separate lodes which carry minable ore in the mines of Zinc Corp. and New Broken Hill Consolidated. The lowest of these and the main ore bodies presently being mined at Zinc Corp. are the lead lodes, consisting of a siliceous fluoritic ore body (No. 3 lens) and overlying it a calcite rhodonitic ore body (No. 2 lens). Each possesses characteristic ratios of lead, silver, and zinc, although the grade within each lens is variable, in places being as high as 30% Pb. There are also low¬grade portions which consist chiefly of poorly mineral¬ized rhodonitic material. No. 3 lens terminates within the Zinc Corp. leases, while No. 2 lens continues into the NBHC leases where it continues to provide a significant proportion of pro¬duction from the mine. Overlying the No. 2 lens are the various zinc-rich ore bodies. In ascending sequence these are lower No. 1 lens, upper No. 1 lens, A lode, and B lode. A typical assay for the small No. 1 lens group ore bodies is 8% lead, 50 g/t silver, and 20% zinc. The `A' lode is a large low-grade ore body in which the higher grade sections assay 4% lead, 30 g/t silver, and 10% zinc; larger quantities of lower grade ma¬terial also occur. B lode, the main zinc lode, typically assays 5% lead, 30 g/t silver, and 17% zinc. B lode increases in size as it extends south, and in NBHC is a major ore body, contributing about 52% of the production. All the lodes pitch south, the average pitch being 0.52 rad (30°) in the Zinc Corp. leases and rather flatter than this in the NBHC leases (see Figs. 1-3). The ground is all competent rock. Production requirements at Zinc Corp. are 900 000 t and 1 200 000 at NBHC.

Jan 1, 1982

By Ken Santini, Ronald R. Bauer, Joseph R. Copeland

Wollastonite, named after W.H. Wollaston, an English chemist and mineralogist, is a calcium metasilicate having the chemical formula CaSiO2. Although it is a unique mineral, wollastonite has a short history of industrial use. Wollastonite is the only naturally occurring, nonmetallic, white acicular mineral. Its acicularity is the main reason for its ascendancy in the late 1970s and 1980s as a reinforcing filler and as a replacement for short-fiber asbestos and short-fiber milled fiberglass. Wollastonite is used in applications ranging from bowling balls to automobile fenders, from ceramic tile to automobile brakes, and from thermal insulation board to protective industrial coatings. The United States is the largest producer of wollastonite, with China, India, and Finland as the other three major sources. The United States, Europe, and the Far East are where most of the production is consumed. Geology Mineralogy and Physical and Chemical Properties Wollastonite, a member of the pyroxenoid mineral group, has the theoretical composition of 48.3% CaO and 51.7% SiO2. Iron, manganese, magnesium, or strontium may substitute for some of the calcium. Commonly associated minerals include garnet, diopside, calcite, quartz, epidote, apatite, and sphene. Some of the physical properties of wollastonite are listed in [Table 1]. Wollastonite is largely chemically inert; however, it can be decomposed in concentrated hydrochloric acid. Some wollastonite will fluoresce under shortwave or longwave ultraviolet light, or both, with colors ranging from pink-orange to yellow-orange and, more rarely, bluish green. In addition, wollastonite may show phos¬phorescence. Because of its unique cleavage properties wollastonite breaks down during crushing and grinding into lath-like or needle-shaped particles of varying acicularity. This particle morphology imparts high strength and is therefore of considerable importance in many of the markets for wollastonite. The acicularity of particles is defined by their length:width, or length:diameter, ratio (i.e., aspect ratio). Individual particles, even the smallest ones, commonly ex¬hibit an aspect ratio of 7:1 or 8:1 and have an average diameter of 3.5 gm. Low aspect ratio products (powder wollastonite, or milled grades) with 3:1 to 5:1 aspect ratios are used as general fillers, in ceramics, and in metallurgical fluxing. High aspect ratio products with ratios of 15:1 to 20:1 are used as functional fillers for rein¬forcing thermoplastic and thermoset polymer compounds and as a replacement for asbestos. The dry brightness and whiteness of wollastonite are also important in determining its suitability for certain filler and ceramic applications. When pure, the mineral is brilliantly white, but impurities may color it cream or gray. Brightness is determined by measuring the reflectance of finely ground powder against a stan¬dard that is assigned a brightness of 100. Magnesium oxide and barium sulfate are the two standards used. G.E. brightness, a term used in North America, refers to brightness measured with a Gen¬eral Electric reflectometer. Commercial wollastonite products usu¬ally have a G.E. brightness ranging from 85 to 95. The Hunter method is also used to measure the brightness of wollastonite. Loss on ignition (LOI), another important property in determining applications for wollastonite, is the amount of volatile matter driven off when the mineral is heated to 1 000°C. Com¬mercial wollastonite products have an ignition loss ranging from 0.5 to 2.0%. Wollastonite's naturally high pH of 9.9 (10% water slurry) is of major importance to its use in the coatings industry. Coatings formulators use milled grades as a pH stabilizer in interior and exterior PVA (polyvinyl acetate) and acrylic latex paints. There are two polymorphs of calcium silicate: wollastonite, a low-temperature form, and pseudowollastonite, a high-temperature form. Conversion of wollastonite to pseudowollastonite occurs at about 1 120°C and results in an increase in the coefficient of expansion and a color change. Pure white wollastonite may change to a cream tint or to various shades of red or brown. This color change is thought to be due to the presence of iron and/or strontium.

Jan 1, 1994

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

The effectiveness of the grouting of saturated, fractured rock in the vicinity of underground workings depends to a great extent on the correct interpretation of the specific hydrogeological conditions at the site. During the planning of grouting operations, calculations of the hydrogeologic co-efficients and of the parameters of the isolating grout curtain(s) and the procedures for forming the curtain(s) must be based on valid hydrogeologic characteristics of the fracture system of the rock within each hydrostratigraphic unit (Anon., 1976). This chapter presents a discussion of the background information that must be collected for these purposes. Details of testing are presented in Chapter 3. The general design of new or the reconstruction of existing underground workings requires a broadly based organization having the expertise to develop a grouting operations plan once it acquires the necessary information. The needed information includes: 1) general information about the hydrogeologic section, 2) the engineering design of the underground workings, 3) the geology of the region, 4) the hydrochemistry of the ground water, 5) the hydraulic properties of the rock, 6) the characterization of all gases present, and 7) characterization of the rock's physical-mechanical properties which is necessary to avoid or control hydrofracturing . The geological documentation portion of the hydrogeological investigation is prepared based on the results of geologic mapping, geophysical surveys and on geo-exploratory operations at a regional and local scale. Geological data from the excavation of other underground workings in adjacent areas of the region is also included. Finally the results of site specific investigations in boreholes constructed at the site of the proposed underground activity are incorporated. A section of the preliminary report entitled "General Information about the Project Site" patterned after the parent STG guidance document on this subject must be com- piled based on the contents of geological reports and on the general construction plan for a new project or the plan for reconstruction of an existing operation. This section of the preliminary report must contain concise information about the conditions under which the planned grouting operations are to be conducted. The information must cover the geo- graphic, administrative, and economic subject areas also. Information about sources of electrical power, heat and water supplies, the presence and characteristics of local construction materials suitable for making grout (particularly clay and cement for the clay-based grout), transportation, and the status of communication lines and other services must be included in this section. Map scales for this section should be on the order of 1:25,000. The technical characterization of planned underground workings is presented in a project report chapter called "General Construction Plan for a New or the Restoration of an Old Operation. " The characterization comprises concise information about the purpose of the workings, the planned depths (for shafts) and the length (for horizontal and inclined workings), the inside and excavated diameters (for shafts), the cross-sectional area (for inclined and horizontal workings), and the face and mouth markers. If excavated underground workings exist in the region, then a concise description of them must be provided. The characterization of all excavations and all operational conditions must be specified. The location within the general structure of the proposed operation is specified. Flow charts must be included in this section that show: 1) the plan for opening the deposit, 2) the plans of the mining operations, and 3) the planned geological sections as they relate to the main workings. These graphical flow charts and appendices are drawn in scales of 1:2000 or 1:5000. 2.1 GEOLOGICAL CHARACTERIZATION The geological characterization of a site of proposed underground workings includes the following principal subject areas: stratigraphy, lithology, structural geology and a characterization of the status of fracturing. 2.1.1 STRATIGRAPHY AND LITHOLOGY A description of the geological section of the rocks to be intersected by the planned underground workings is complied on the basis of: 1) research data derived from core extracted from exploratory boreholes and monitoring boreholes, 2) the documentation and mapping of all natural out- crops in the region, 3) the data from all existing under-ground workings in the regions where operations are being canied out already, along with data from adjacent regions, 4) the data from surface and borehole geophysical investigations (standard and specially developed logging methods, and 5) regional stratigraphy. The geological description of the rock must express the maximum detail possible. The description provides the name of the rock, the granulometric composition (for fragmented rock), the composition and type of stratification, the drilling characteristics of the rock, a visual estimate of porosity and density for each stratum identified, the karstic

Jan 1, 1993

By W. –E. Marx, M. Faber, R. Pollak, I. Astberg

INTRODUCTION Essentially three strategies are pursued in parallel for protecting the underground working force in mining against the consequences of methane ignitions which might lead to explosions: • Sufficient ventilation for avoiding the formation of explosive methane-air mixtures. • Avoiding sources of ignition caused possibly by electrical equipment or friction. • Structural explosion protection for limiting explosion propagation to the farthest-going extent for the case that methane is fired because of failing of the above preventive measures taken. The corresponding methods taken in mechanized roadheadings with point-attack roadheaders in German coalmining are subject of this paper. VENTILATION In ventilation of roadheading sites the following is to be considered in respect of methane control: • Sufficient air flows by volume being cycled to the heading face for diluting released methane to safe concentrations. • Sufficient local air velocities for effective mixing of methane and air in the whole heading site, in particular directly on the heading face and underneath the roof. Furthermore, ventilation needs to be laid out in a way that cut- ting does not result in health affecting dust exposure at any working place. Accordingly, the ventilation systems shown on Fig. 1 including primary and, if necessary, secondary auxiliary ventilation air ducts as well as dedusters, are used in roadheadings. This equipment is run according to the following principles: • If the roadheader is at standstill, the heading face is ventilated by the free air jet from the blowing duct closest to the face. In this operation phase the deduster is shut-down. • During cutting, the exhaust of the blowing air duct is automatically closed, and the deduster is taken into operation. The in- coming fresh air is now locally discharged via a lateral orifice, or distributed by Coanda-type ducts from the primary or secondary ventilation air duct. In this way, a stable dust wall not disturbed by the jet of the air duct is formed in front of the heading face. Only the ventilation air stream taken in by suction of the close-to-face deduster is passed through the inby zone of the heading face (over approx. 20 m). The clean exhaust of the deduster is dischayed to the roadway in some distance outby. Immediately at the heading face, methane accumulations possibly released by cutting are diluted by compressed-air jets from air-movers located near the roof which assure more rapid motion of the air without distuhing the dust wall. • In the zone of overlapping of primary and secondary duct and the deduster - that zone possibly is poorly ventilated during cutting - sufficient local air speed is assured by Coanda-type ducts or additional ventilation jets for avoiding methane accumulation. The regulations of the inspectorate specifies geometrical conditions to be met with these configurations of ventilation equipment. In-seam roadways are ventilated at rates > 0,5 m/s. For dedusters, present-day ventilation air flows are of 600-800 m+/ min. Ventilation monitoring and control is assured by units for measuring methane and ventilation air flows. If pre-set thresholds are exceeded or fallen short of, these measuring units shut-down - dependent on potential hazard - any not intrinsically safe electrical equipment in the roadheading, either totally or in limited zones. In addition, CO-measuring units near the roadway en- trance serve for early detection of mine fires. In spite of this ventilation layout and the equipment used and in spite of the water spray systems on the cutting head (discussed later in this paper) several cases of methane ignition and subsequent burning off (however, no explosions) have been recorded in the recent past. Therefore, the above-described ventilation is nowadays complemented in many cases by machine-mounted ventilation equipment for the immediate surroundings of the cutting head. Ventilation of the surroundings of the cutting head Investigations carried out subsequent to the above-mentioned gas ignitions have shown that the high ventilation air speed necessary for rapid dilution of methane could not permanently be arrived at in the following zones: • The direct zone of cutting, in particular during sumping. • The floor zone above the debris, predominantly below the cut- ting head. Accordingly, VOEST ALPINE-Bergtechnik, a producer of mining machinery, in cooperation with DMT developed for point- alttack roadheaders a novel machine-mounted ventilation equipment, the Jet-Block" (AERO SAFE Jet-Block) shown on Fig. 2. This system allows to ventilate the above-mentioned critical zones with high flow speeds. One AS Jet-Block is mounted on each side of the boom approx. 0.5 m behind the cutting head. The blocks discharge compressed-air jets with high energy from 10 nozzles per block into the hazardous zones (Fig. 3). Two nozzles of each jet block are tangentially oriented toward the upper circumference of the cutting head, the others are arranged fan-like with orientation to the heading face underneath the cutting head and the debris. The nozzle array of the blocks as well as the com-

Jan 1, 1997

By Bruce H. Mason

Lightweight aggregates are classified as those minerals, natural rock materials, rock-like products, and byproducts of manufacturing processes that are used as bulk fillers in lightweight structural concrete, concrete building blocks, precast structural units, road surfacing materials, plaster aggregates, and insulating fill. Other uses include architectural wall covers, suspended ceilings, soil conditioners, and other agricultural uses (Mason, 1989). One unusual agricultural use recently reported was for bedding on a crocodile farm on the south coast of Natal. It seems that female crocodiles became upset because there was no vermiculite in which to lay their eggs due to the late arrival of a consignment of vermiculite. The vermiculite was being used as a substitute for river sand to nest crocodile eggs (Griffiths, 1989). The various lightweight aggregate materials may be classified into four groups as follows: 1. Natural Lightweight Aggregate Materials-prepared by crushing and sizing natural rock materials, such as pumice, scoria, tuff, breccia, and volcanic cinders. 2. Manufactured Structural Lightweight Aggregates-prepared by pyroprocessing shale, clay, or slate in rotary kilns or on traveling grate sintering machines. 3. Byproduct Lightweight Aggregates-prepared by crushing and sizing foamed and granulated slag, cinders, and coke breeze. This group of lightweight materials, with the exception of slag, has become less significant and will not be discussed in this chapter. Other lightweight aggregates are beginning to appear in the marketplace. Some are the result of manufacturing processes on industrial byproducts and/or waste products, such as fly ash and incinerator residue (Bigner, 1992). 4. Manufactured Insulating Ultralightweight Aggregates-pre- pared by pyroprocessing ground vermiculite, perlite, and diatomite. The application of diatomite in brick and in extruded and aggregate forms has declined in recent years (Kadey, 1983) and will not be considered in this chapter. Lightweight aggregates for structural concrete and concrete masonry units, according to the American Society for Testing and Materials (1986) consist of "aggregates of low density used to produce lightweight concrete, including: pumice, scoria, volcanic cinders, tuff and diatomite; expanded shale, clay, slate, perlite, vermiculite or slag; and end products of coal or coal combustion" (Mason, 1989). Lightweight aggregates are distinguished from other mineral aggregate materials by their lighter unit weights-1 281 to 1 602 kg/m3 or less for loosely packed lightweight aggregates-as com- pared to 2 002 kg/m3 or more for crushed stone, sand, gravel, air-cooled slag, and other common mineral aggregates. The specific gravities of these lightweight aggregates are also used to classify them for use in concrete; they range from the extremely light materials, used for insulative and non-structural concrete to expanded shale, clay, slate, and slags used for structural concrete. Vermiculite and perlite are sometimes referred to as super light- weights. Insulating concrete (non-load bearing) weighing as little as 240 kg/m3 can be made with these aggregates. Natural aggregates such as pumice, pumicite, and scoria are normally made into concrete weighing about 400 to 480 kg/m3. Expanded shale, clay, and slate produced in a rotary kiln make structural concrete ranging from 1 201 to 1 842 kg/m3. Expanded shale and clay produced by sintering make concrete that ranges from 1 442 to 1 922 kg/m3. Air-cooled slag produces concrete ranging from 2 162 to 2 403 kg/m3. The low density lightweight aggregates are used primarily for insulating purposes. They have relatively low compressive strength. The middle density aggregates are used for insulation and for fill. Those at the upper end develop excellent compressive strength and are used in many structural applications (Anon., 1971). In recent years, the production of lightweight aggregates in the United States for all uses (including insulation and construction materials) has averaged about 19 Mtpy (Table 1). The production and consumption of lightweight aggregate is closely tied to the fortunes of the construction industry and the need to construct energy efficient buildings. Recent years have seen the production of lightweight aggregate approach the levels of the early 1970s. The period of 1979 to 1981 was a low point in production levels. The low levels of production were caused by significantly higher plant energy costs due to the large amounts of fuel and energy used in the production of most types of lightweight aggregate. Lower levels of construction activity also affect production levels. This chapter focuses on the use of natural lightweight aggregates and manufactured insulating ultralightweight aggregates, and on both the production and use of manufactured structural light- weight aggregates. Some additional comments are made on byproduct materials. Separate commodity chapters on vermiculite, perlite, pumice, and volcanic cinders are elsewhere in the volume. Discussion of the commodities covered elsewhere in the book will be limited to observations on their utilization and relation to other lightweight aggregate materials.

Jan 1, 1994

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

5.1 TECHNOLOGY FOR INJECTING GROUT INTO BOREHOLES The grouting of fractured rock aquifers is carried out in separate stopes' whose size is determined by the hydro¬stratigraphy and the hydrogeologic properties of the rocks. The number of grouting stopes required depends on the number of intervals in the hydrogeological section. Each grouting stope can include one or several aquifers. Com¬bining several water-producing strata into one stope is pos¬sible only when the values of their permeability coefficients are very similar. In order to achieve efficiency, as many aquifers as possible are included in each stope. In practice, the length of a grouting stope in a borehole usually ranges within 10 to 40 m, but occasionally up to 100 m or more. When conducting grouting operations through boreholes drilled from the face of underground workings, the stope's size can be determined by the thick¬ness of each fractured zone (fault) or by the length of the hole if the entire rock is fractured. Injecting grout into separate intervals of fractured rock can be done from the bottom to the top with ascending stopes or from the top to the bottom with descending stopes. Injection from bottom to top is always preferred. Injection also can be accomplished in separate fractured permeable horizons in the section. The STG integrated grouting method is based on the use of packer and injection equipment and technological schemes that facilitate injection in highly isolated fractured aquifers. For this purpose STG has developed a standard series of packers. 5.1.1 INJECTING GROUT THROUGH BOREHOLES DRILLED FROM THE SURFACE Grout can be injected according to one of four techno¬logical schemes depending on the depth of the fractured aquifers, their hydrogeologic properties and the availability of grout injection equipment if the area is remote (Fig. 43). In the first and most preferred technological scheme (1), the grout is injected from the bottom up with ascending stopes. When this scheme is employed, grouting holes are drilled to the entire projected depth based on the field investigations injection tubes is installed above the top of the first hydrostratigraphic unit to be grouted. This scheme al¬lows the grout to be injected only into the permeable stra¬tum or strata isolated by the packer near the bottom of the hole; it precludes the need to clean or redrill the hole before injecting grout into the next overlying stope. This ascending stope scheme is preferred because of its simplicity and ef¬ficiency. It is used in most of the grouting of saturated fractured rock for the purpose of isolating aquifers while excavating deep mine shafts in the USSR. The second technological scheme (II in Fig. 43) injects the grout in sequence from top to bottom with descending stopes. Injection under this scheme is accomplished through the hermetically-sealed mouth of the hole. It is most advis¬able to grout with descending stopes in saturated fractured rock wherein the aquifers constitute thick networks of frac¬tures. Under these conditions it is practically impossible to drill a grout hole to the entire depth of the underground workings due to the large production rate of ground water. In conducting isolation operations with descending stopes, additional work (relative to ascending stopes) is necessary to drill out the grout in order to create and inject the next stope. As a result, the total amount of drilling required to drill out grout can even exceed the volume of drilling of original rock. Consequently it is advisable to use this inefficient scheme only for injecting grout into holes that contain prolific frac¬tured aquifers at depths less than 150 to 200 m. The third and fourth technological schemes (III and IV in Fig. 43) are used in the selective grouting of separate aquifers. For example, these methods would be applied when a grout hole is drilled to its entire depth, wherein grouting of the upper aquifers must be done first so that the sinking of the first shaft can begin prior to deeper grouting. The third technological scheme is used for the selective injection of grout into an aquifer with a thickness of up to 5 m. For this work the DAU-1-2 removable dual packer is installed. It is lowered into the borehole on an injection pipeline. For aquifers with thicknesses of more then 5 m, the selective injection of the grout is carried out according to the fourth technological scheme; a combination of the DAU-2 hydrodynamic plug and the DAU-1 packing device is used. In order to implement this injection scheme, the DAU-2 hydro-dynamic plug is installed at the bottom of the fractured aquifer and the DAU-1 packer is installed at the top of the aquifer.

Jan 1, 1993

By Warren D. MacEvoy

INTRODUCTION Good ventilation bulkheads (stoppings) can be obtained only by conscientiously adhering to four basic principles: carefully selecting the site for the bulkhead, adequately preparing the site, properly constructing and sealing the bulkhead, and checking the results with a smoke tube before leaving the area. If all these guidelines are followed during each ventilation project, a poor job will result only from the use of improper materials and cannot be blamed on carelessness or oversight. SITE SELECTION A perfectly constructed bulkhead in the wrong location can be completely worthless. Conceivably, a bulk-head in the wrong location could cause dangerous ventilation conditions that are difficult to identify and isolate. If the suggestions herein are followed, a considerable amount of work, time, and material may be saved. Perhaps the single most important step is always to select the exact site for a bulkhead before starting construction. Although the latitude in selecting a site may be limited by instructions or operating conditions, it is common to find poor bulkheads built within a few meters (feet) of a site that would have been far superior but was overlooked during the planning process. The ideal site for a bulkhead should have as many of the following features and conditions as possible: 1) To provide a solid bulkhead, the selected site should be in firm and unbroken ground. 2) To minimize the amount of work and material required and to minimize the danger of later damage to the bulkhead, the site should have as small a cross section as possible. 3) The bulkhead site should not be encumbered by interferences such as pipes, wires, ditches, wire roof supports, trash, muck piles, etc. 4) The walls and timbers of the site should be free from oil, grease, or tar that would inhibit adhesion of the sealant used on the bulkhead. 5) Unless special precautions are taken, the site should not have water seepage, standing water on the floor, or water-carrying ditches. 6) To expedite construction of the bulkhead, the site should have reasonable access to transportation, supplies, communication facilities, and compressed-air lines. 7) The site should have a reasonably level floor, allowing direction of the bulkhead door swing to be reversed at a later time if so desired. In a steeply sloping location, it may be possible to open the door in one direction only. SITE PREPARATION Once a suitable site has been selected for the bulk¬head, it must be prepared properly. All site preparation work should be completed before starting any construction work on the stopping itself. Unfavorable site conditions can be identified during the course of thepreliminary site preparation, saving time, effort, and materials that otherwise might be wasted. The follow¬ing preparation steps help assure the construction of a good stopping: 1) If wire mesh has been used at the site, a strip 457 mm (18 in.) wide should be cut and removed from the walls and roof at the selected location. 2) The exposed strip of rock should be barred down thoroughly to provide a smooth surface. 3) All obstructing materials should be removed, including old timbers, pipes, rockbolts, wires, etc. If a conduit must cross the bulkhead area, it should be located or relocated away from the floor, walls, and roof to allow a 6.28-rad (360°) seal around the juncture between the conduit and the bulkhead structure. 4) All loose muck should be cleaned from the site and a trench about 152 mm (6 in.) deep should be dug from wall to wall in the floor. The trench must be wide enough to accommodate the entire bulkhead, including the posts. 5) The rock surfaces of the walls and roof should be cleaned with a wire brush to remove as much loose surface material as possible. Thorough cleaning promotes adhesion of the bulkhead sealant to the surfaces of the walls and roof, thus promoting an airtight seal. 6) It is quite difficult to seal landing mats that cross a bulkhead. If such a crossing cannot be avoided, the bulkhead should be placed between the end and the first hole in the steel so at least one side of the bulk-head can be sealed easily and completely. That may have to be done prior to actual construction if the end is located on the opposite face from the seal coating of the bulkhead. BULKHEAD CONSTRUCTION The three principal considerations in bulkhead con¬struction are the type of bulkhead, materials to be used, and construction method to be employed. For many projects, the bulkheads are specified by the requesting agency, with no latitude for independent choice of the type, materials, or construction method. In such cases, any deviations from the specifications, for any reason, must be approved in advance by the proper department or by the project senior ventilation engineer. Types and Materials When a choice of type, materials, or method is allowed, consideration should be given to factors such as cost, required useful life, proximity to blasting concus¬sions, availability of materials, direction of permissible air leakage, and degree of airtightness required. Other factors to be considered include the amount of time available, accessibility of transportation, potential for interference with operations or production, ambient water conditions, availability of connections to com¬pressed-air lines, etc. Despite the multitude of factors to be considered, most stoppings can be analyzed easily and the proper choices can be made without much difficulty. The four common classes of bulkheads utilized in

Jan 1, 1982

By Lok W. Home

INTRODUCTION Today, raise drilling (boring) has become the stan¬dard method of raising throughout the Western world. It is estimated that the majority of raise footage com¬pleted in 1977 was accomplished by the raise drilling method. Extensive replacement of the drill and shoot system by mechanized raise drilling has occurred be¬cause users have recognized one or more of the follow¬ing advantages. In a direct cost per meter comparison, mechanized raise drilling is economically more attractive than drilling and shooting. Mechanized raise drilling offers improved personnel safety over the drilling and shooting method. Mechanized raise drilling is a faster and more predictable method for excavation than is drilling and shooting. The raise cost per unit of length becomes less as the raise length increases, offering greater flexibility in the overall design of mines. Mechanized raise drilling provides a substantial reduction in labor requirements. 6) Mechanized raise drilling improves the working environment. 7) Mechanized raise drilling improves the rock in¬tegrity of round, unblasted raises. Increasing acceptance of the raise drilling technique is enhanced by continuing equipment development. Today, a wide range of machines and capabilities is available, machine drive systems have been improved, reliability has been increased, and cutter technology is being improved quite rapidly. Table 1 lists the four major manufacturers of raise drilling machines, cutters, and reamers, together with the approximate number of machines that each company had in operation at the end of 1977. Although Reed and Smith are major suppliers of reamers and cutters, they do not manufacture the drilling machines. Four or five other relatively minor suppliers have built machines, reamers, and cutters for raise drilling, but they did not have a substantial num¬ber of machines in the field at the end of 1977. MACHINES FOR MECHANIZED RAISE DRILLING There are three principal arrangements for making machine-bored raises. They include the use of standard raise drills, reversible raise drills, and blind-hole or box¬hole raise drills. Standard Raise Drills The standard raise drilling machine is set up on one mine level or on the surface, and a pilot hole is drilled from that level or surface to a level below. When the pilot hole has been drilled, the raise then is reamed back from the lower level to the upper level or the surface. This arrangement is illustrated in Fig. 1. Reversible Raise Drills The reversible raise drilling machine is set up on a lower level of the mine, and a pilot hole is drilled up¬ward to a higher level. The raise then is reamed back from the upper level to the lower level. This arrange¬ment is illustrated in Fig. 2. The machines used for this type of raise drilling normally are capable of being used as standard raise drills. Blind-Hole or Boxhole Raise Drills The blind-hole or boxhole raise drilling machine is set up on a lower level of the mine, and a full diameter raise is bored to a higher level without the use of a pilot hole. This type of machine is illustrated in Fig. 3. Drill Recommendations If access to upper and lower mine levels is equal, the standard raise drill should be used. The complexity of the equipment and associated handling procedures for boxhole drilling far outweigh any mechanical advantage or reduction in the number of operations, provided that access to an upper level is available. If there is good access to a lower level but only limited access to an upper level, the use of a reversible raise drill should be considered. If there is no access to an upper level, a blind-hole or boxhole raise drill is the only practical mechanized system. MACHINE SELECTION There are six major considerations in selecting a machine for raise drilling. These include operating range, cutter loading, machine structure and main¬tenance, drill string configuration, drive system, and automated controls. Operating Range As shown in Table 2 and Figs. 4 through 6, raise drilling machines are available for a wide range of hole sizes. Standard raise drill ratings range from raises of 0.9 m (3 ft) diam and 120 m (400 ft) length up to raises of 3.5 m (12 ft) diam and 900 m (3000 ft) length. Boxhole raise drilling machines are available for raises ranging from 0.9 m (3 ft) to 2 m (6 ft) diam with lengths up to 90 m (300 ft). Table 2 lists, among other criteria, the torque and thrust of typical machines; torque and thrust are the main criteria for analyzing the capacity of a machine. The raise sizes shown in Table 2 assume a condition of medium hard to hard rock formations and optimum cutter loading. The recommended raise size for a given machine can be exceeded when raising softer rock formations or when the operator is willing to accept a lower level of performance; the performance is related directly to the 1093

Jan 1, 1982

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

6.1 PRINCIPLES OF PLANNING GROUTING OPERATIONS FOR SHAFTS, DRIFTS AND TUNNELS The planning of integrated grouting of vertical, inclined or horizontal underground workings in saturated, fractured rock includes the following phases: 1. Planning for the collection and analysis of geological information concerning the characteristics of the struc¬tural deformation(s) that produced the details of the frac¬ture system being studied. 2. Planning for the acquisition of initial hydrogeological information produced by the flowmetric and hydrody¬namic investigations (described in Chapters 2 and 3). 3. Computing the initial engineering calculations of the ex¬pected or planned dimensions of the isolation curtains for each water-producing stratum (fracture zone) and the expected areal spreading pattern of the grout in each of these water-producing horizons. The selection of the lo¬cations, the initial optimal number of holes and the vol¬ume of grout needed to create each isolation curtain also are included in this phase. 4. The selection or development of a grout formula for the specific hydrogeological conditions that exist at the site. 5. Identification of a resource (deposit) required to produce the grouting clay. The planning of grouting operations must be supported by the acquisition of all necessary initial data, which in¬cludes: 1. The geographic location and administrative responsibil¬ity (ownership and regulatory constraints) of the mineral deposit and any peculiar features of the area where ex¬cavation of the underground workings is planned. The surface relief, the hydrographic network of the region, and the local climatic conditions must be evaluated prior to planning for buildings and waste discharges if any. 2. The economic status of the region, the available trans¬portation, the nature of the energy base, the water and heat supply sources, the location of construction mate¬rials suitable for making the grout all must be deter¬mined. 3. The geological and hydrogeological characterization of the work site must be accomplished. 4. The technical features of the shaft(s) and underground works must be characterized. 5. The organization's principal initial design plan features and the construction sequence of all buildings and struc¬tures must be developed. These must be correlated with the organization and schedule for excavating the under¬ground works. The technical characterization of a shaft or of other un¬derground workings must contain information about the purpose of the workings, the design depth, the planned excavation procedures along with cross-sections, and the parameters of the lining, if a lining is required. The geology of the section to be penetrated must include geological data obtained from existing exploration activities at the mineral deposit and conclusions about the shaft ex¬cavation conditions obtained from drilling technical moni¬toring holes and exploration drillholes. The interpreted stratigraphic regime, including the thickness, strike and dip of each stratum, must be shown in cross-section. The tec¬tonics, including all geologic structures, must be presented by cross-sections. The bedding orientations must be used to identify anticlines, synclines, monoclines and faults. The lithologic composition and the drillability of the rock com¬prising the section must be examined. The character of the rock's fracture pattern must be presented. Finally the per¬meability, porosity and storage characteristics of each aqui¬fer must be determined by in situ hydraulic property testing (injection tests or pumping tests, as described in Chapter 3). The fracture pattern of the indurated rock is presented in reports derived from geo-exploration borehole studies and from studies of outcrops on the surface and in all nearby underground workings. The extent of each fracture zone is mapped in three dimensions. The distribution of fracture apertures, the nature of the fractures according to each litho¬logic stratum, and the relationship of bedding planes to each fracture system are all determined. Acoustic logs, downhole cameras, caliper logs, oriented core, and any other appro¬priate geophysical tools available are used to construct the data base for this purpose. The hydrogeological characterization of the geologic section to be penetrated by the workings is based on the analysis of hydrogeological data collected during the drill¬ing and testing of technical monitoring holes and from in¬formation on ground water inflow collected during the ex¬cavation of nearby underground workings, if such workings exist. The hydrogeologic characterization must contain the following information: 1. The identification of all aquifers, their thicknesses and their locations in the section, 2. The distribution of hydrostatic head in the ground water in three dimensions, 3. The coefficient of saturated hydraulic conductivity (per¬ meability) of each aquifer, 4. The porosity and storativity of each aquifer, and 5. The projected rate of water inflow into the shaft, drift or tunnel for each aquifer prior to grouting. This step is used to identify the zones to be grouted.

Jan 1, 1993

By Tony Walsh

Australian Stock Exchange Limited (ASX) -ASX is ranked 9th in the Morgan Stanley Capital International (MSCI) Index – only 2.02% of that Index. -1520 companies – 400+ mining & mining exploration. -Over 50% of Australians own shares in ASX listed co.s -Disclosure standards play a vital part in providing investors with confidence, thereby building depth and liquidity in the market.

Jan 1, 2003

By Joaquin Ruiz

Although Mexico is known for its important base-and precious-metal deposits, and lately oil, some industrial minerals play an important role in the Mexican economy. Sulfur, the most important industrial mineral produced, accounts for approximately 7.5 percent of the total value of mineral production. This commodity is followed by fluorspar, which accounts for 6.4 percent. Predominantly, the fluorspar deposits are very high grade (75 percent CaF) and are contained in lower Cretaceous limestone in the states of Coahuila and San Luis Potosi. Base-metal deposits in the area of Parral (notably San Francisco del Oro), however, have become important fluorite producers in their own right. (San Francisco del Oro has 12 percent fluorite gangue.) Salt from the Guerrero Negro solar-evaporation ponds in Baja California account for 6 percent of the total shares; phosphate from Baja California also accounts for 6 percent. Barite from man to deposits in Lower Cretaceous limestone in Coahuila and exhalative deposits in the Paleozoic section of Sonora account for 1.0 percent. In addition, Mexico is an important producer of gypsum, graphite, and celestite. Although graphite and celestite are not large-scale operations and thus do not contribute much to the Mexican economy, the Mexican graphite and celestite deposits are among the most important in the world; Mexico is one of the world's few sources of celestite. The celestite is in the form of large, high-grade mantos in Cretaceous limestones in northern Mexico.

Jan 1, 1987

By John C. Balla

The Troy deposit is located in northwestern Montana, approximately 17 mi south of Troy. The deposit is a stratiform, strata- bound, copper-silver deposit which occurs in the Revett formation of the Proterozoic Belt Supergroup. The Revett formation is approximately 1730 ft thick in the vicinity of the Troy deposit, and is divisible into three units. The upper Revett, approximately 250 ft thick, is the host for the orebody. The Revett represents a deltaic environment. The Troy deposit was discovered by Bear Creek Mining Company in 1 963 during the course of a reconnaissance stream sediment sampling program that was designed to discover strata-bound copper deposits. The deposit is approximately 7500 ft long, 1800 ft wide, and averages 60 ft thick. It is bounded on the north by outcrop, on the east by a post-mineral fault, on the south by diminishing grade and post-mineral faults, and on the west by diminishing grade. Ore reserves are 64 million tons with a grade of 0.76% Cu and 1.58 oz Ag per ton. The mine is presently operating at a production rate of 4.2 million oz of silver and 20,000 tons of copper per year.

Jan 1, 2013

By A. L. Holmes

Since the 1981 record of 54 mm tons of US fertilizer consumption, grain surpluses in the US have brought about massive acreage reduction programs, decreased fertilizer consumption, and resulting price pressures caused by excess capacity. The strong US dollar that reduced export sales of both grain and fertilizer, and new offshore manufacturing capacity using US engineering/contracting know-how have combined to put US producers of phosphate fertilizers in uncompetitive postures. While recent events suggest the industry may be growing out of this capacity problem, the cost of US phosphates will continue to be disadvantaged by such external factors as taxes, environ- mental regulations, and high power costs. There is no reason to expect help from the government in the form of reduced environmental regulation and taxation. In order to survive, phosphate fertilizer producers will have to develop technologies and adopt strategies that result in world competitive production costs. Current responses to the need to reduce costs include reduced energy costs through cogeneration, innovative reclamation techniques that create valuable land, and in- creased raw material recovery achieved by operator proficiency. While new technology offers long-term benefits for the producer, present efforts are concentrated on more mundane management efforts to reduce costs. Major attention is being given to the problems of labor utilization and productivity. Raw materials suppliers are being asked to renegotiate contracts which more appropriately reflect conditions in an industry operating at about 50% of capacity. Achievements in these and other areas will allow the producers to continue operating until longer term benefits materialize. `

Jan 1, 1986

By B. Hearn, O&apos

Falconbridge, Ltd. operates six mines (the strathcona, Fraser, Lockerby, Onaping, East and Kidd Creek) in the Sudbury and Timmins districts of northern Ontario (Fig. 1). These mines vary in depth to approximately 1500m and experience a wide variety of ground conditions. Problems of greatest concern include rockbursting, backfill pillar stability, wall and pillar stability, and the use of cable anchors. A study was initiated approximately two years ago to identify how numerical models could aid in the examination and possible understanding or solution of these problems. This paper discusses the use of numerical models at Falconbridge and describes a case history of their use in examination of rockbursting at the Strathcona Mine.

Jan 1, 1988