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By Charles T. Steuart, Richard B. Berg

Tripoli and the related mineral commodities such as micro- crystalline silica have been mined for more than 100 years for their abrasive properties. Although abrasive and buffing compound markets are still very important, within the last 15 years the filler and extender markets in paint, plastics, rubber, adhesives, and sealants have increased substantially. Tripoli or microcrystalline silica consists almost entirely of very small quartz crystals, many less than one micrometer in length. Material mined from different districts differs in crystal shape, grain size, and texture of the rock, all of which influence markets. Most US deposits are now mined by surface methods, and both air floated and micronized products are marketed. Deposits of tripoli now mined in the United States occur in chert-bearing Paleozoic limestone in the central part of the country with producing districts in southern Illinois, central Arkansas, and eastern Oklahoma. Although deposits within each district are confined to specific formations and extend laterally within those formations, individual bodies form minable deposits that are typically several hectares in areal extent. Tripoli is white to cream to rose and characterized by high porosity and ease of disaggregation. DEFINITIONS Tripoli In the United States, tripoli was first used to describe the fine-grained, easily disaggregated material from Seneca, MO, be- cause of its similarity to a rock from the Tripoli region of North Africa (Hovey, 1894). The North African rock is actually diatomaceous earth, a material that is similar in appearance to the rock from Seneca, but is of entirely different origin having formed by the accumulation of siliceous remains of microscopic marine or fresh- water animals. Tripoli is best defined as a very fine-grained, generally porous rock that consists of microcrystalline quartz, typically formed by the alteration of a chert-bearing limestone. Tripolite A term used to describe a rock from the vicinity of Tripoli in North Africa which is diatomaceous earth (Quirk and Bates, 1978) Microcrystalline Silica Microcrystalline silica is the same material as tripoli, but the distinction between the use of these two names is dictated largely by convention and markets. Material produced from southern I1linois deposits and used in white pigment and filler applications is generally referred to as microcrystalline silica, whereas that used in abrasive applications, both from the Illinois district and from other states, is commonly called tripoli. Amorphous Silica Amorphous silica, a term formerly used to describe the material produced from the deposits in southern Illinois, is now replaced by the term microcrystalline silica. Amorphous silica came into use when even optical methods for the identification of very fine- grained quartz were not widely available and the Illinois product, composed of quartz grains too small to be seen with the unaided eye, was thought to consist of amorphous silica. The Illinois material is clearly crystalline quartz, as shown by X-ray diffraction analysis and scanning electron microscopy (Fig. 1). Novaculite Although originally used to describe a rock suitable for the manufacture of whetstones, novaculite is now defined more generally " - as a homogeneous, mostly white or light colored rock, translucent on thin edges, with a waxy or dull luster, and almost entirely composed of microcrystalline quartz" (Steuart et al., 1983). The more compact rock mined in central Arkansas from the Arkansas Novaculite is referred to as novaculite, whereas the more porous rock is referred to as tripoli. Rottenstone The commodity rottenstone is sometimes included within the general mineral commodity category of tripoli. Rottenstone is mined in Northumberland County in eastern Pennsvlvania and formed by the weathering of a siliceous shale of Devonian age (Faill, 1979, Berkheiser, private communication, 1991). This material is used as a filler and extender, but is apparently unlike tripoli both in origin and physical properties. Spiculite Spiculite is a rock consisting of siliceous sponge spicules having formed by the removal by solution of the carbonate matrix of a spicule-bearing limestone. Spiculite has been mined in Texas and, because it resembles tripoli in several aspects, is included in the discussion of deposits.

Jan 1, 1994

By Wm. J. McCracken

HISTORICAL REVIEW OF BOARD OPERATIONS The Ontario Worker's Compensation Board was established in law enacted by the legislature of the Province of Ontario in 1915. It was designed to pay insurance benefits to injured workers, and at the same time to protect employers from legal suit. It was based upon an enquiry system rather than an adversary system such as that used in the courts process. Initially, the system was designed to pay compensation benefits and subsequently, to pay for the cost of medical treatment and pensions for disability and disease resultant from the effects of traumatic injury. In 1947, the Act was changed to include industrial or occupational generated diseases, not specifically related to traumatology. Such occupational diseases were therefore accepted and benefits paid subsequent to that date. As will be discussed in several minutes, even today the vast preponderance of compensation claims with the Ontario Board continues to be related to the effects of trauma. HISTORICAL REVIEW OF EXPOSURE TO RADON GAS DECAY PRODUCTS In some areas of Ontario, especially in Northern Ontario, there is a natural leaching of radon gas from the underlying rock formation. This constitutes very low levels of radon gas decay product radiation exposure to those persons coming in contact and inhaling these substances. This paper however is designed to discuss the occupational generated types of radon gas exposures. For many years dating back to the 1930's, partially refined ores were being shipped from Northern Canada to a refinery located at Port Hope, Ontario, still in operation and currently operated by Eldorado Nuclear Limited of Canada. Initially, the purpose for the operation was extraction of radium to be sold on world markets for medical treatment purposes. With the advent of World War II, this market collapsed. Subsequent to World War II, the availability of other sources of radiation for medical radio-therapy generally replaced the requirements for radium. During World War II, a new market opened up for the Port Hope refinery however as work into nuclear chain reactions and the development of the atomic bomb identified the need for uranium and enriched uranium. During the period of operations where radium was being extracted at the Port Hope refinery, it is now known that an identifiable radon gas hazard did exist. This hazard disappeared when the production line for extraction of radium ceased operations. In 1954, uranium mining operations opened up in Ontario at two locations, Bancroft and Elliot Lake. At the peak of operations, 16 mines were operational and 11,000 workers were employed in these mining operations. A high level of mining activity continued over a 10 year interval with the Bancroft Mines closing permanently in 1964 following a 10 year life of operation. The other mines in Elliot Lake closed about the same time with the exception of two uranium mine operations which have continued to be operational up to the present time. By 1965, due to a dramatic drop in world demand for uranium, the total work force had fallen to 1/10 of the peak work force, and approximately 1,300 workers remained in employment. It is of interest to note that one significant difference in the work environment between Elliot Lake and Bancroft was the high silica content of the Elliot Lake ore. This gave rise to a number of cases of silicosis developing in relatively short intervals of time in the Elliot Lake miner population. No cases of silicosis were identified from the Bancroft operations. Based upon the experience in investigating and evaluating actual cases of lung cancer in the uranium miners over the years, the medical staff at the Ontario Board also developed the impression that radiation levels were much higher in the Bancroft operations, especially in the earlier years of operation, than at Elliot Lake. This resulted in accumulation of higher levels of Working Level Months (WLM), usually over a shorter exposure interval in many of the cases. This aspect will be further evaluated in this presentation. Subsequent to 1965, the work force remained quite static in numbers until 1975. At that time, there began to develop an increase in the work force, and this increase is continuing at a moderate rate up to the present. INITIAL METHOD OF HANDLING LUNG CANCER CLAIMS The first lung cancer claims in Ontario from uranium mining operations were accepted on the perceived cause-effect relationship. This relationship was based upon the data from the Colorado observations and the Czechoslovakia data. Initially, a series of regression equations on mortality were developed and used to estimate the effect of exposure to low cumulative doses of radon daughters as it might relate to the frequency of occurrence of lung cancer at any particular cumulative exposure level. A probability of cancer being radiation induced as against it being caused from other factors was calculated. This method was discontinued subsequent to 1972 due to problems encountered in carrying out this complex evaluation. Thereafter, each case was dealt with on an individual basis, being based upon whether or not the tumour was of the oat cell type, a cumulative exposure in excess of 120 WLM; latency periods in excess of 10 years, commencement of mining prior to

Jan 1, 1981

By William S. Pintz, William A. Vogely, Charles J. Johnson

This chapter on the role of minerals in developing economies is oriented toward economic policy issues. The discussion and analysis cut across the technical, economic, legal, and policy fields. The goal of this chapter is to pro- vide relevant background information important to individuals involved with the minerals sector in developing countries. Where possible, trends are shown to provide a basis for understanding the evolving role of minerals in developing economies. This chapter is divided into the following five sections: (1) characteristics of developing countries (which presents selected statistical comparisons); (2) government objectives in developing countries; (3) government policy formulation in developing countries; (4) determinants of future trends, and (5) a summary of the chapter's major points combined with the authors' view. CHARACTERISTICS OF DEVELOPING COUNTRIES There are 155 nations in the world categorized as "developing countries." In total these nations represent about one third of the value of the world's mineral production, three quarters of the world's population, and over half of the world's land surface area. The group of so-called developing countries is quite diverse and there are exceptions to almost any general observation about the group as a whole. Most definitions of the term "developing country" are based on levels of per capita income. However, if per capita income is the sole criterion used in categorization, then several oil-exporting countries would jump from the category of developing country to the category of developed country.a Obviously, this results in a poor fit with respect to other measures. In general, the majority of developing countries can be characterized by relatively low per capita income, high population growth rates, high levels of unemployment, low levels of education, short average life ex- pectancies, income distribution skewed toward the highest 20% of the population, and a large share of exports represented by primary commodities. b For the purposes of this chapter a slightly modified version of the widely used World Bank system of classification (shown in Table 1.2.1) is used. As summarized in Table 1.2.1, developing countries are divided into three main groups based on estimated gross national product (GNP) per capita. It should be noted that the high- income, oil-exporting countries have characteristics of both developing and industrialized countries; however they are included for most statistical comparisons within the group of developing countries. Of the 155 developing countries in the world, approximately half produce minerals other than common building materials, and only 32 account for more than five percent of world pro-

Jan 1, 1985

By K. S. Venkataraman

The physical properties of clay-water systems depend on the complicated system of forces between the clay particles themselves, and between the clay particles and the ions in the liquid phase. The kind and distribution of ions in, on, and between the clay particles and the size and the shape of the particles are the basic factors determining the macroscopic behavior of clay-water systems. Understanding the system requires a knowledge of the nature of the clay particles, their size, structure, composition, and surface properties, and of the manner in which they interact with ions [and molecules] in the surrounding liquid [or other medium]. The validity of Professor Brindley's words (Brindley, 1958), written three decades ago in the context of making pottery, whitewares, and electrical porcelains, transcends time, and the basic message is perhaps all the more important in the considerably expanded use of ceramics for structural, thermal, tribological, electronic, and other applications. Silicon carbide, silicon nitride, and sialons have been studied in the last two decades for high- temperature structural and tribological applications, particularly for using in internal combustion engines. Titanates, zirconates and niobates of barium, strontium and lead, have high dielectric constants, and are extensively used in the formulations for making capacitors. Hexagonal ferrites (molecular formula MO.6Fe2O3) are in use for making permanent magnets for fabricating miniature motors, and for assembling loud speakers, particle accelerators etc. Cubic ferrites such as magnesium-zinc ferrite and nickel-zinc ferrite are used as transformer cores, and for other high-frequency applications. In this context, Richerson's recent book (Richerson, 1984) on the general scope of traditional and technical ceramics is a good starting point for an overview of contemporary ceramics technology. Glasses are a whole class of amorphous materials used widely as sintering aids, and for making glass-bonded ceramics and glass-ceramic composites. Composites are yet another burgeoning field where two or more particulate components are used for improving the performance of ceramics. For all these applications, the inorganic starting materials are almost always submicron and near-micron powders. Understanding the powders' physicochemical properties, and their surface chemical interactions with the surrounding liquid/gaseous medium is-necessary for making reliable ceramic parts at competitive prices. Even though ceramics science and engineering has attained its separate identity in universities and the industry, ceramists themselves would concede that ceramics science is a cross-disciplinary field, having incorporated and assimilated within itself many principles from several apparently disjointed disciplines. Principles of material science, graduate-level physics and chemistry, polymer science, surface and colloid chemistry, transport phenomena, particle technology, unit operations commonly used in chemical engineering and mineral processing, and statistics and applied mathematics are integral part of any ceramics curriculum in universities. Added to this is the fact that all bench-scale successes in making ceramic parts are to be scaled-up for larger throughput operations. Understanding and applying process engineering principles of comminution, classification, drying, calcination, etc. then becomes essential. CERAMIC FORMING: Despite the diversity of the materials and processes, conceptually, the steps involved in making ceramic parts have remained the same over several decades: The different components for making the pan (usually one or more powders plus other forming and sintering additives) are proportioned and mixed thoroughly, and the well-mixed formulations are consolidated into desirable shapes known as "green bodies." Usually binders such as wax, clay, organic polymers and surfactants, whether dispersed or dissolved in a suitable liquid are used during mixing the batch for giving strength for the green bodies. In the dried green state, the inorganic powders typically occupy only 55 to 60% of the bulk volume of the body, depending on the particle size distributions of the powders and the forming history, with mostly inter- particle voids accounting for the rest of the void volume. SINTERING: The formed bodies are then fired in high- temperatures kilns/furnaces during which the parts are exposed to a predetermined temperature profile, and "soaked" for a certain duration at the final high temperatures, typically between 1200 K and 1900 K, and then cooled to room temperature. The gaseous atmosphere in the furnace is controlled (oxidizing, reducing, or inert) when necessary. During the initial stages of firing, volatile liquids evaporate, and during the intermediate temperatures between 400 and 600 K, the the organic polymeric additives pyrolize and oxidize into water vapor, CO, C02, and other gases. At still high temperature, the glasses, when present, soften, and simultaneously, the ceramic particles rearrange into a network of grains with definite grain boundaries so as to reduce the total interfacial free

Jan 1, 1990

By Martin H. Stanczyk

Under the Solid Waste Disposal Act, the Bureau of Mines initiated several accelerated in-house research programs to develop new and improved methods of proper and economic solid waste disposal. The program being conducted at the Bureau' College Park Metallurgy Research Center is concerned with eliminating the current waste of valuable materials in municipal refuse by developing low-cost methods for reclaiming and recycling the metal and mineral values contained in incinerator residues. Each year the United States must collect and dispose of in some manner or other, about 150 to 200 million tons of municipal solid wastes. The monetary output for this service ranks third among public services, exceeded only by the costs for schools and roads based on population growth projections and materials use patterns, the quantity of refuse generated by the early 1980's may approach 300 million tons annually. Currant estimates indicate that only 10 to 20 percent of the total municipal refuse generated is disposed of by incineration. If all refuse were properly incinerated the residues would contain more than 11 million tons of iron and nearly a million tons of mixed nonferrous metals including aluminum, copper, lend, zinc, and tin.

Jan 1, 1969

By James R. Arnold, Cindy S. Jones, Michael F. Gleason, John O. Marsden, John G. Mansanti

INTRODUCTION The Chimney Creek Gold Mine (Gold Fields Operating Co. - Chimney Creek) is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. The operation is a wholly owned subsidiary of Gold Fields Mining Corporation, the North American branch of Consolidated Gold Fields PLC, London, England. The plant started up in November, 1987, less than three years after discovery of the orebody and three months ahead of schedule. Ore is mined in an open pit and is processed by combined dump leaching and milling techniques for gold and silver recovery. The mine is set to produce approximately 150.000 ounces of gold and 50,000 ounces of silver per year over a 12 year life at current reserve estimations. The mine was designed and constructed at a cost of $79.3 million with engineering and construction services provided by Davy McKee Corporation, San Ramon, California. Key Gold Fields operating staff were involved in the design of the facility from the start of the project: The Mine Manager, Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist were all involved full time on the project within 5 months of the first ore discovery. Emphasis was directed at optimizing operating efficiency and in particular minimizing labor costs in the plant. It was recognized that a high level of instrumentation and control would be required to achieve this. The risk associated with the instrumentation and control systems implemented was to be minimized by using equipment and systems that had been proven in industry while utilizing the most cost effective, state-of-the-art technology available. The reliability of the overall control system was considered to be critical in view of the cost of downtime associated with the gold extraction plant. BRIEF PROCESS DESCRIPTION The dump leaching process treats approximately 1.2 million tons per year of low grade ore at an average grade of 0.035 oz/ton. Run of mine material is dumped on a lined leach pad and weak cyanide solution is applied by drip irrigation. Pregnant solution run off is pumped to carbon columns in the milling plant for gold recovery and the barren solution returned to the dump leach circuit. Average gold recovery is 60%. This process has little instrumentation and control associated with it. The milling operation treats 700,000 tons annually of higher grade ore (0.200 oz/ton initially, dropping to an average of 0.135 oz/ton after first two years). Recovery is directly related to head grade (fixed tail assay effect) and currently averages 96%. A single pass through a jaw crusher reduces run of mine ore to minus 12 inches. The ore is stockpiled and reclaimed by loader for grinding in a two-stage milling circuit consisting of a SAG mill and ball mill, the latter in closed circuit with hydrocyclones. Cyanide and lime are added into the SAG mill to start dissolution of gold as early as possible in the circuit. The ground product leaves the milling circuit at approximately 78% minus 200 mesh and is fed to an unique "double thickener" leaching-recovery circuit. This circuit has been discussed in detail in a paper by J. G. Mansanti et a1 (1). Two thickeners are arranged in counter- current configuration with three leach tanks. Overflow solution from the first thickener is treated by carbon-in-columns (CIC) for gold recovery with 85% of the soluble gold recovered onto this carbon. Underflow slurry from this thickener is pumped to the leach tanks, with a total retention time of 12 hours, and then gravitates to the No. 2 thickener. Overflow solution from the second thickener is used as a wash in the first thickener. Underflow slurry from the second thickener is treated in a carbon-in- pulp (CIP) scavenging circuit to recover the remaining 15% dissolved gold. Gold-loaded carbon from both the dump leach and milling circuits is stripped in batches using the Zadra hot caustic- cyanide elution process. Gold (and silver) is recovered from the hot strip solution by precipitation with zinc dust and the product recovered on Funda pressure filters. The precipitate is retorted to remove any mercury and then smelted into buttons. The buttons (approximately 80% gold, 15% silver) are shipped to an independent refiner in Salt Lake City, Utah, for further treatment.

Jan 1, 1990

By S. L. Bessinger

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

Jan 1, 1992

By L. G. Stolarczyk

In-mine Continuous Wave (CH) Radio Imaging method (RIM) instrumentation has the capability of measuring EM wave propagation constants in layered formations. The RIM method is based upon the fact that seams of coal, trona, potash, quartzite, gilsonite and oil bearing sands form natural waveguides for low loss transmission of electromagnetic (EM) waves. The waveguide forms whenever the seam is bounded by more conducting claystone, mudstone, shale or in the case of oil/gas traps, by a floor of saline water. Disturbances caused by faults, shear zones, paleochannels, sills and dikes significantly alter the propagation constant. The analytical values of the EM wave propagation constants depend on the physical and electrical parameters of the seam. The values are presented in this paper in a graphical form to assist in the interpretation of RIM tomographic images. This paper compares in-mine geologic mapping obtained during mining to tomographic images of the geologic disturbance zones in a Central Utah coal mine.

Jan 1, 1989

By C. A. Rawlins

INTRODUCTION Radcont is a program designed by the author of this paper for the industry to use as an instrument for radiation contamination evaluation and ventilation planning system. Radiation in mines are associated with the mining of gold and gold bearing minerals, as uranium and thorium is incorporated in the mining of these minerals. Radiation contamination in South African mines is not a new concept as it was investigated by the Chamber of Mines in the early 1960's and found not to be hazardous at the time. Since some of our mines export scrap metal to customers abroad, it came to light (1991) that some of the scrap metal was radioactive. The authority that oversees the nuclear aspects in South Africa is the Council for Nuclear Safety (CNS). They investigated these matters and found that the mines needed further information regarding radioactive material and the handling of these contaminated materials. As the various mines were licensed (with various conditions incorporated) thereafter, the mines had to do their own investigations as to what extent their properties (Surface and underground) were radioactively contaminated. Some mines were found to be highly contaminated over the years of operation and controlling conditions were installed and measures installed to reduce the contamination levels. One of the conditions when issuing a licence by the Council for Nuclear Safety (CNS), is that a screening survey be carried out to determine the radiation exposure levels and corrective action to be taken if necessary. These surveys must be done by a person trained in the required procedures for such a survey. The person must also measure the risk correctly and assess the results properly. In such a survey, the internal and external exposure levels must be determined to assess the total exposure of persons working in those conditions and take appropriate action if necessary. When doing such a survey, hundreds and more likely, thou- sands of data points are recorded. In order to assess the data recorded, various integrated and difficult calculations need to be made, and takes up enormous amounts of time. (This excludes the interpretation of the results ) The following explanation of the program shows the different parts of such a survey assessment calculations to be done. The paper details the program layout and the different sub- sections within the primary program. It must be stated that the program, as with any other program, is as accurate as the data inserted into the data base. The program and details thereof are given under the following headings: 1. TOTAL EFFECTIVE DOSAGE WITH REGARDS TO: • GME required gravimetric results obtained (mg/m3) • Thick layer or total contamination measured (Bq/m2) • Dry condition surveys with dust loads taken as a Standard l0mg/m3 • Wet conditions survey with dust loads taken as l mg/m3 • Airborne long lived alpha and beta activities as determined by analysis in Bg/m3 • LTD (Thermoluminescent Dosimeter). Results as obtained from the SABS (South African Buro of Standards) are recorded in this section for each month of the year for each individual worker. An average dose is then determined at the end of the year. • Bucket measurements as recorded. • Smear samples (Loose contamination). As determined by Electra or by analysis • Occupational factors for Metallurgical and Engineering occupations in and around the Metallurgical facilities of your mine. • All underground dosage determination and calculations. (Radon and Thoron) 2. INFORMATION REQUIRED WHEN PROGRAM IS INITIALISED: As the program is started, it opens up on the contents page. Here there are various options to choose from, but one is cautioned as a beginner in operating the program, not to perform any tasks before carefully reading these instructions. Firstly, one must go to the 'Information required" pushbutton. Press this button. The information required page is shown where the cursor can be moved to the block where one can enter the specific mines name. To enter a mines name, put the cursor in the block provided and just insert the mines name with the normal keyboard keys and press the enter button on the computer keyboard. To enter the other information required such as Alpha and Beta instrument efficiency, ALI (Annual limit of intake) and probe area, one can either press the 'Data required" button for a dialog box information or enter it manually by just putting the cursor in the block provided and entering as did above. In order to insert all the required information for the pro- gram to calculate the information required, one must proceed further by entering the area names surveyed in the spaces provided. There are 20 spaces to enter 20 different areas surveyed. One must further also provide the amount of days worked in each area (i.8. 250) in the block provided. The de- fault is 250 days. There are also standard information given in the information data page such as breathing rate (1,2 m31h), 8 hours worked per day, 5 days per week and 50 weeks per

Jan 1, 1997

By S. A. Tims

INTRODUCTION The Falconbridge mine ore body extends about 1.6 km (1 mile) in length and the deepest developed ore is on the 6050 level below surface. The ore zone varies in width from a few inches to over 30 m (100 ft) and the average width is 4.9 m (16 ft). Access levels are driven in the ore at 53.3-m (175-ft) intervals. The principal method of mining is overhand longitudinal cut-¬and-fill. Prior to 1962 timber square-set stoping as a secondary extraction method was used for about 15% of the total production. Undercut-and-fill was intro¬duced at Falconbridge in 1962 as a potential replace¬ment for the square-set method in heavy ground. The undercut-and-fill method was developed by Inco in the 1950s, its principal application being to transverse pillar mining. Falconbridge made modifications to this method. A feature of the mine is the No. 1 flat fault which dips 0.79 rad (45°) towards the northeast. The main characteristic of the fault is the presence of large swells of ore directly under the plane of the fault. The ground under the fault area is highly fractured and associated with massive sulfides. In the past, the ore under the fault was recovered, with difficulty, by either tight cut¬and-fill or square-set stoping. In the 1970s these meth¬ods were supplanted to a large degree by the under-cut-¬and-fill method. An advantage of the current undercut-and-fill method which uses cemented fill compared to the cut-and-fill and square-set methods is the reduction of dilution due to better control of the walls. At Falconbridge mine, it is estimated that the grade of ore produced by undercut¬and-fill is improved by approximately 10% over other methods. Where undercut-and-fill is used in very weak ground, a much greater improvement in grade can be expected. Table I shows mining production for 1974. The undercut-and-fill method was first used at Falconbridge during 1962. The first longitudinal stope was prepared for undercutting by laying down laminated beams the length of the stope and installing a lagging mat floor on top of the beams. Unconsolidated tailings fill was poured on top of the mat floor. As the cut ad¬vanced under the floor, heavy posts were placed under the laminated beams at 1.8-m (6-ft) intervals. During 1966, a radical change was made to the method when tailings fill, consolidated with portland cement, replaced the unconsolidated fill. This development eliminated the laminated beams and heavy mat floor and greatly im¬proved the stability of the stope. This system, with minor variations, is currently used at Falconbridge mine. APPLICATION The undercut-and-fill method is used to mine in¬competent ground, sills or floor pillars under mined-out levels, or a block of ore isolated between levels. It is occasionally used to advantage in sequencing produc¬tion from various mining blocks. This is done by mining a block of ore cut-and-fill method and at the same time mining the ore block directly underneath by the under¬cut-and-fill method. The undercut-and-fill mill holes at Falconbridge are either boreholes, stripped timbered raises, or steel mill holes. Boreholes and rock raises tend to slough in heavy and broken ground which increases dilution when sloughing exceeds the ore width outline and also in¬creases the difficulty of moving down to start the new cut. For example, in one installation, a 1.2-m (4-ft) diam borehole sloughed to a size of 3.7 x 5.5 m (12 x 18 ft). The undercut-and-fill method usually requires a mill hole extending from the level below the ore to the top horizon of the ore block. The customary methods of providing a mill hole are: 1) A borehole is driven from level to level through the ore block and a chute installed on the bottom level (Fig. 1). 2) An existing raise is used as a mill hole. If the raise is timbered, a steel mill hole is installed inside the timber and tailings fill poured around the steel mill hole (Fig. 1). 3) An existing steel mill hole, situated at one end of a mined-out stope, is used as the mill hole for an ad¬jacent undercut-and-fill ore block. The mill hole posi¬tion is determined when planning the mining sequence of the first stope (Fig. 2).

Jan 1, 1982

By Ivor Hawkes

INTRODUCTION High energy breaking is an alternative to using ex¬plosives in underground secondary breaking operations. It also is a means of upgrading conventional hand-held breakers, manual sledge-hammer breaking, and scaling bar operations. Major areas of application are in sec¬ondary breaking over grizzlies and at drawpoints. Other applications include breaking down ripping lips in longwall seam mining, scaling in stopes and rooms, general demolition work, and roadway maintenance. There is considerable interest in high-energy impact breakers for use in primary ore breaking, but, as of 1977, all such applications have been only experimental (duToit, 1973; Joughin, 1976; Wayment and Grantmyre, 1976). EQUIPMENT Essentially, a high-energy impact hammer is a boom¬mounted pneumatically or hydraulically actuated breaker. The machine basically consists of a piston that oscillates in a housing and impacts the end of a tool or moil thrust against the rock. The force applied to the rock primarily depends upon the impact energy of the piston-the higher the impact or blow energy, the greater the force and, thus, the greater the rock break¬age. Among drill and breaker designers, a common expression for blow energy is "force of blow." Hand-held breakers are limited to blow energies of about 140 J (100 ft-lb), because the operator is unable to handle heavier machines efficiently or to absorb the recoil energy resulting from higher blow energies. How¬ever, these restrictions do not apply to boom mounted breakers; machines with blow energies on the order of 4000 J (3000 ft-lb) and higher are available commer¬cially for underground use. There is considerable evi¬dence to show that increasing the blow energy also in¬creases the efficiency of the breaking operation; i.e., more rock is broken per unit of energy expended (Grantmyre and Hawkes, 1975). Thus, there is a trend to higher blow-energy machines, particularly where high¬strength rocks are to be broken. In relation to rock breaking, the blow rate of boom¬mounted impact breakers is not as important as it is for rock drills. This is because the breaker must be moved over the work surface between blows. The blow rate is governed eventually by the power supply, and typical blow rates range between 200 and 600 blows per minute. As a general rule, light blow-energy machines have higher blow rates than heavier machines. Table 1 lists most of the boom-mounted impact breakers that were available commercially during 1977, and it gives details of the blow energies and machine weights. Restrictions are placed on the blow energy by the machine weight and size, and by the strength of the boom. Typically, boom-mounted impact hammers have a blow-energy to mass ratio of about 1.5, with lower values for lighter machines and higher values for heavier machines. In addition to supporting the hammer weight, the boom also has to absorb the recoil energy of the blow, which can be on the order of 1400 J (1000 ft-lb) for large hammers operating in a horizontal mode. Interesting exceptions to the general run of impactors are the Joy HEFTI hydraulic hammers. In these machines, the piston impacts onto a fluid cushion that is positioned between the piston and the impact tool. This approach allows very high piston velocities, over 30 m/s (100 fps), to be used without the risk of break¬ing the piston or impact tool. Steel on steel impacts must be limited to impact velocities of about 10 m/s (35 fps) due to the high impact stresses generated; thus, increased blow energies can be achieved only by increas¬ing the piston size. The Joy 514 HEFTI®, listed in Table 1, has a blow energy of 27 100 J (20,000 ft-lb), but, as of 1977, the machine has been used underground only on an experimental basis. Using a fluid cushion between the piston and the impact tool allows the use of light pistons, reducing the overall machine weight. The recoil energy, which must be absorbed by the boom for a given blow energy, is directly proportional to the piston to machine mass ratio, and operating with light pistons provides an addi¬tional benefit in reducing the requisite boom size. Both pneumatic and hydraulic hammers are avail¬able commercially. Although hydraulic hammers are a relatively recent development, they already outnumber the pneumatic machines in use. There are many reasons

Jan 1, 1982

By P. R. ZETTWOOG

HISTORICAL BACKGROUND A program in France to develop personal [a] dosimeters has been initiated 1974. The patent on which is based the present device was obtained in 1972 * . From 1972 to 1974, the possibilities of applying certain ionograph track detectors to the spectrodosimetry of radon daughters was explored. The first prototype were produced in 1974. It took four years (from 1974 to 1978) to produce an autonomous dosimeter whose components has a sufficient life span, especially for the turbine motor unit. Qualification in the laboratory was obtained in 1977. In 1978 it was obtained in the mine for technology (autonomy of 12 hours and a life span of more than one year) and in 1980 for monitoring. 300 dosimeters have been tested in underground mines all together. Indispensable peripheral equipment were also developed from 1976 to 1980 : calibration devices, equipment to prepare and develop the films, read out systems. The concept of an "Integrated System of Individual Dosimetry" (ISID) based on a personal [a] dosimeter measuring exposure to radon daughters, thoron daughters, ore dust and external irradiation doses was proposed at the end of 1980. Since January 1st 1981, ISID is used on a routine basis in some french mines, situated in remote area, and appears to be very competitive with the ambiant dosimetry. The latest version of the dosimeter is produced in mass series since June 1981 and should equip all french mines in 1982. DESCRIPTION OF THE INSTRUMENTATION DEVELOPMENT OF THE DOSIMETER MEASURING HEAD The measuring head is based on the use of ionographic film to detect a tracks. In fact, the measuring head is a spectrodosimeter which measures separately over the period of exposure: - the potential [a] energy inhaled due to the decay of Po 218, Po 214, and Po 212 ; - the number of Rn 222 atoms inhaled ; - the inhaled total [a] activity of the five long-lived emitters present in the ore dust. The contribution to the total inhalable potential [a] energy of these various radionuclidesin a typical underground mine is studied in Appendix I. The measuring head described in detail in Appendix II, is able to satisfy all the implications made in the ICRP recommendations. Appendix III deals with the use of this measuring head in the cases where the equilibrium factor is lower than 0.1. This situation occur in open-pit mines where account must be taken of the Rn 222 contribution, which is no longer negligible in relation to that of its daughters. CURRENT DOSIMETER PERFORMANCES Table I shows the characteristics of the latest dosimeter. Appendix IV should be consulted concerning qualification of the dosimeter in the laboratory and in mines, technological development which finally produced the [a] dosimeter and its peripheral equipment, and technical presentation of the ISID (Integrated System of Individual Dosimetry) based on the concept of a multirisk personal dosimeter. Data on the installation and operating costs of such a dosimeter, which would seem to be competitive, are also given in this Appendix. ADVANTAGES OF PERSONAL DOSIMETRY AS COMPARED TO AREA MONITORING The results of the first eight months of experiments carried out under real conditions in an underground mine site are given in detail in reference 8. Area monitoring : The monthly exposure per worker to inhaled Rn 222 was determined from the knowledge of time spent in various areas of the mine and for the different mining operations, as well as from numerous and systematic sampling of the Rn 222 concentration in all work places. Personal dosimetry : The exposure to potential energy from radon daughters was measured by an [a] dosimeter developed by the CEA and worn by each of the 160 miners during eight months. In this way 160 x 8 pairs of monthly individual exposure values have been obtained which can be statistically studied. This test was decisive for us because it proved that the [a] dosimeter was technically sound (very few defects over one year for 160 dosimeters) and especially that personal monitoring devices were superior to area monitoring devices. The following conclusions can be drawn 1. The exposure distribution obtained by personal dosimetry is log-normal. This is true for the results on the whole as well as for groups of results relating to certain explanatory variables. See fig. la, 1b, 1c. 2. The exposure distribution obtained by area monitoring does not correspond to any type of distribution. If the results of personal monitoring are taken as a reference, area monitoring tends to underestimate the high exposures and overestimate the low exposures.See fig. 2. 3. [a]-energy exposures are underestimated when calculated from radon exposures and the equilibrium factor found in the considered mines. This is due to episodes or to zones of high radon concentrations not registered

Jan 1, 1981

By Theodore M. Gurr, James J. Bartels

Phosphate produced fertilizers provide the phosphorous nutrients required by plant life to sustain significant growth, thus improving the production of food for the world's population. Phosphorous replenishment of soil currently can be obtained efficiently by no other means than by direct application of phosphate fertilizer to the soil. Phosphate products are also utilized in animal feeds, detergent, and various industrial processes. Thirty-three countries are presently mining phosphate rock for the production of fertilizers, and its process byproducts. The major phosphate rock producers are the United States, the former Soviet Union, Morocco, and China [(Table 1)]. Phosphate rock is generally defined as a rock material that contains phosphate minerals which are sufficient for commercial usage. Phosphate minerals are found in sedimentary igneous and metamorphic rocks. Economic resources of phosphates are primarily developed from sedimentary rock sources. The apatite mineral family comprises the majority of phosphate constituents from both sedimentary and igneous rocks. Igneous phosphate rocks are generally composed of fluroapatite whereas sedimentary rocks are generally composed of carbonate fluroapatite, or chlorapatite. Phosphate rock is mined by a multitude of processes, including surface and underground mining. The mined rock requires mechanical and chemical processing to liberate the phosphate for utilization as fertilizer. Approximately 85% of the world's phosphate production uses sulfuric acid, with the remaining processes utilizing nitric and phosphoric acids. Ammonia is also introduced in the manufacture of liquid fertilizer, ammonium phosphate and ammonium polyphosphate. GEOLOGY Mineralogy and Chemical Properties Phosphate most commonly is derived from the mineral apatite, which is chemically described as CA5(F,Cl,OH)(PO4)3. Apatite is also broken up into the composite apatite minerals which are fluroapatite Ca5F(PO4)3, chlorapatite Ca5Cl (PO4)3 and hydroxyl-apatite Ca5(OH)(PO4)3. Carbonate CO3 can substitute for PO, forming carbonate apatite, called francolite. The guano phosphate (bird excrement) mineralogy commonly occurs as brushite CaHP04.2H20, monetite CaHPO4, whitlockite Ca3(PO4), and dahlite Ca10(PO4)6(OH)2. Apatite minerals form from igneous sources where they are developed deep in the earth's crust; and the cooling processes are slow, producing long, hexagonal prismatic to tabular crystals. The termination of the crystal can be basal plans or pyramids. In some instances, the crystals are bipyramids. Faster cooling, or sedimentary derived apatites are more commonly cryptocrystalline. Collophane is a name given to cryptocrystalline apatite found in phosphate rock in fossil form. Apatite minerals are relatively soft and are used to describe the 5 hardness on the Mohs scale. Cleavage is poor and develops along the C(0001). The color of apatite minerals varies from colorless, to violet, and blue, but are predominantly green or brown; luster is vitreous to subresinous. Collophane's physical appearance is often opaline, dense, with colloform structure and sometimes concretionary, nodular or pulverulent. Phosphate Bearing Materials Origins Phosphate deposits of igneous and metamorphic origins have been well defined. Sedimentary origins have been defined, but with a greater degree of uncertainty. Research in the 1980s proved out some of these processes through studies of real time deposition in Chile, Peru, and Australia. Igneous intrusive alkali rock and associated contact metamorphic rocks, provide approximately 20% of the world's phosphate. Fluorine containing apatite minerals are the most common materials containing phosphorous. Fluorapatite occurs in most igneous rocks. Fluorapatite and fluorine hydroxylapatite, together with carbonate varieties of these, are important members of the group. Where an essentially pure chlorapatite, carbonate apatite, and hydroxylapatite are rare and restricted in occurrence, the fluorine containing types occur in most all igneous rocks as early formed accessory mineral, usually in microscopic crystals, and may occur as extremely large bodies as magmatic segregations from alkalic igneous rocks. Apatite is also found crystallized in pegmatitic faces of both acidic and basic types of igneous rocks. More specifically, apatite is associated with magnetite deposits, in hydrothermal veins, especially those formed at relatively high temperatures, and in veins of the Alpine type. Apatite is common in both regionally and contact metamorphosed rocks, especially in the crystalline limestones where it is associated with sphene, zircon, pyroxine amphibole, spinel vesuvianite, phlogopite, talc, chloride schists, and as a contact metamorphic mineral.

Jan 1, 1994

By Victor E. Archer

INTRODUCTION Man is exposed to many agents which induce mutations in germ cells and/or cancer at work, at play, and at home. In this total mix of mutagenic and carcinogenic agents, how important are radon and its daughters? Before man moved into caves and other permanent dwellings, the principal mutagenic and carcinogenic agent to which he was exposed was natural background radiation--cosmic rays, radium and potassium-40 in his food, plus gamma rays and radon from the soil and rocks. When man moved into caves, captured fire, and began to preserve and store foods, his exposure to carcinogens and mutagens took a quantum leap. Carcinogens and mutagens appear to act in the same way, that is, by altering the DNA or nuclear proteins of cells. Most mutagens are carcinogens, and vice versa, so when I say mutagens from here on, I will be referring to both. The relationship of the two is emphasized by the fact that administration of a carcinogen to a group of animals not only increases cancer rates among the exposed animals, but also among their progeny (Tomatis 1979). Environmental Mutagens Smoke from man's fires, overheated foods, and foods preserved by smoking, resulted in ingestion and inhalation of many polycyclic aromatic hydrocarbons--many of which are mutagens. Caves and houses with tight windows and doors tend to collect the radon which is constantly emanating out of soil, rocks and concrete, so man's exposure to the radon daughter component of background radiation increased several fold. Preserving food by salting or pickling with material that contained nitrites and nitrates led to increased ingestion of nitrosamines, which are potent mutagens. When his grains and other foods were stored in slightly damp rooms, fungi or mold would grow on them. Several of these fungi are now known to produce very potent mutagens. The best known of these is aflatoxin B (Ramachandra 1979). It may seem strange that a living organism would produce a mutagen. One might think that it would scramble its own genetic heritage. The reason it does not is that it produces the mutagen in an inactive form. It can be activated only by an animal's enzyme systems after being eaten. When man moved into cities, the collective smoke from wood and coal fires further increased his exposure. That particular smoke has now mostly disappeared, but has been replaced by smoke from automobiles and industry. When man moved into the age of technology, his exposure to mutagens again increased dramatically. Many mutagenic chemicals, from benzene and beta naphthylamine to a long array of pesticides and tobacco products have been added to our environment. Excess deaths from cancer are now being observed among chemists in most industrialized nations. Mutagens are even found in much of our wine, beer, and whiskey (Keller 1980). Some of the chemical mutagens were widely used in food or in other commercial products before their potential was discovered. Striking examples of this is the original butter coloring agent and the polychlorinated biphenyls that have been widely used in brake fluids and electrical transformers. Large quantities of them have been discarded or disposed of in a careless manner--in such a way that many of them have contaminated our food, our ground water and air (Landrigan 1981). In this nation, with the help of several recent laws, we were just beginning to get control of the industrial chemical mutagens. With the relaxing of these laws that is currently going on, it appears that it will be many more years before we really bring chemical mutagens under control. Many nations have yet to come to grips with this problem. On top of this massive array of chemical mutagens we have now added radiation from many artificial sources. For most of us this means medical X-ray and fallout from nuclear weapons testing. Ionizing radiation is one of the most potent mutagens, so it has caught the public eye, and its contribution cannot be ignored. Fortunately, by the time we started using radioactive materials in quantity with the Manhattan Project, we had experience with radium and X-ray (some of it bad); we knew enough radiobiology and enough about methods of radiation protection so that most nuclear laboratories have had a phenomenal record of radiation safety. Radiation is one new technology with great potential for harm that has not exhibited that potential except for a few isolated situations like that of radium dial painters, uranium miners and atomic bomb victims. Uranium miners slipped into this list almost by accident. We could have protected our uranium miners just as well as we did the workers in nuclear laboratories; but we failed to do so. Why didn't we? The reason is simple. The Atomic Energy Commission was charged with protecting the health of their workers. They did not wait for a pile of bodies before they introduced controls. Congress appropriated the money, and taxpayers were willing to pay for the protection against radiation. Miners unfortunately did not work for the Atomic Energy Commission. Although mine operators were ignorant about radiation, the key item was that in the 1950s nobody was willing to pay the extra costs of adequate ventilation to control the high levels of radon and radon daughters in uranium mines. Control was not achieved until new laws and regulations were passed which made it compulsory. BIRTH DEFECTS AND CANCER

Jan 1, 1981

By M. J. Howes, C. A. Nixon

INTRODUCTION A safe heat stress control strategy for an underground mine has three elements: Application of an environmental measure which reflects physiological strain with sufficient accuracy for the range of conditions encountered underground. Acceptance of a functional relationship between the environ- mental measure and human performance which is used to optimise the environmental conditions achievable with either ventilation or ventilation and refrigeration. A management control strategy based on the environmental measure which is designed to ensure that work in environments where excessive physiological strain may occur is prevented and corrective action is initiated. The environmental measure that reflects physiological strain is the link between the three elements and, since the turn of the century, the discussion of the merits of various indices has been prolific. One problem in selecting a suitable measure or index is the ease with which it can be physically obtained relative to accurately reflecting the physiological strain. For example, wet bulb temperature is simple to measure and, for a particular mining sys- tem, it may adequately represent physiological strain, however, it would not necessarily provide the same relatively safe measure in a different mining system. The acceptance of a measure which can be universally applied has been confounded by both development and predisposition. That is not to say that there is only one "correct" measure and all others are unsuitable. It is self evident that if the application of a particular index has resulted in adequate control, then that mea- sure is correct for that situation. However, an understanding of the limitations is necessary to ensure that adequate control is maintained as mining conditions change. Almost 100 years after the question of heat stress in mines started to be dealt with in a collective manner, an analysis of the available information is leading towards a general strategy to control this problem. In the paper, the developments in heat stress assessment are briefly examined and followed since the earliest published observations on the effect of heat in mines (Haldane, 1905), efforts to determine a relationship between an environmental measure and human performance are reviewed and summarised and the benefits of control strategies such as acclimatisation and shortened shifts are discussed as they relate to Mount Isa Mines. The results of testing the prototype air cooling power instrument are discussed and a heat stress control strategy outlined. HEAT STRESS AND AIR COOLING POWER The operation of the human engine is analogous to other engines where the conversion of chemical energy from the oxidation of fuel to useful mechanical energy is not 100% efficient. In a diesel engine it is about 33% and in a human engine less than 20% resulting in at least five times as much heat produced by the meta- bolic process as useful work done. Metabolic energy production is related to the rate at which oxygen is consumed and is about 340 W for each litre of oxygen per minute. Using measured oxygen consumption and an average body surface area of 2.0 m2, the approximate metabolic energy production associated with different mining tasks is (Morrison et al. 1968):- • Rest, 50 W/m2 • Light work, 75 to 125 W/m2 (machine, LHD or drill jumbo operators) • Medium work, 125 to 175 W/m2 (airleg drilling, light construction work) • Hard work, 175 to 275 W/m2 (barring down, building bulkheads and timbering) • Very hard work, over 275 W/m2 (shovelling rock) Heat balance is achieved when the rate of producing heat (the metabolic heat production rate) is equal to the rate at which the body can reject heat mainly through radiation, convection and evaporation. Heat exchange between the lungs and the air in- haled and exhaled is normally less than 5% of the total and there- fore usually ignored. Any heat not rejected to the surroundings will cause an increase in body core temperature. Since heat stress is related to the balance between the body and the surrounding thermal environment, the main parameters required to be known when determining acceptable conditions are those associated with the heat production and transfer mechanisms. These can be summarised as follows: Metabolic heat production rates (M - W) Skin surface area (A3) (and effects of clothing) Dry bulb temperature (t[ ]) Radiant temperature (t[ ]) Air velocity (V) Air pressure (P) Air vapour pressure (e [ ]) The rate of heat transfer to or from the environment depends on the equilibrium skin temperature t, and the sweat rate S,. These in turn depend on the response of the body to the imposed heat stress and the effect of thermoregulation (Stewart, 1981). Thermoregulation The body contains temperature sensitive structures which send impulses to the brain at a rate depending on the temperature. Both hot and cold signals can be differentiated and the thermoregulatory response ahivated according to which signal pre- dominates. If "cold" signals are dominant, body heat loss is re-

Jan 1, 1997

By M. J. Hughes

Kaiser Steel Corporation's Eagle Mountain operations are located in Southern California at a distance of about 130 miles east of the Corporation's steel plant at Fontana, California, and at a distance of about 225 miles from the nearest harbors which are located at Long Beach and Los Angeles, California Although the Eagle Mountain ore bodies are strategically located with respect to the steel plant and the harbor facilities, they are not of a very high grade and 'the geologic nature of the deposit makes it very difficult to mine and to concentrate. The iron ores occur as partial replacement of older zones mainly within dolomites and to a lesser extent within quartzites and quartz monzonites. The alteration was extremely variable and generally resulted in the formation of pyritic magnetic which, in part of the ore body, was leached to remove the pyrite and change the magnetite to hematite. Although the overall ore body is large, a major portion of it contains sulfur in the range of 1.0% to 5.0%. Obviously, this is too high a sulfur content for direct use in blast furnace feed.

Jan 1, 1968

By S. J. Bluhm

INTRODUCTION Lesotho is a mountainous area of southern Africa from which water is to be exported in an extensive tunnel system, to industrial regions inland. The related tunnelling project has involved a num- ber of drives using tunnel boring machines [TBMs] to excavate about 100 km of 5 m diameter water tunnels [von Glehn and Bluhm, 1995). This paper describes the ventilation and cooling of some of the tunnel drives from both the operational and design points-of-view with a particular focus on heat generation. There were many common features in all of the drives but this paper is focused mainly on the Hlotse drive which was 18,4 km long. The drives were ventilated using forced ventilation systems to provide appropriate air flow throughout the tunnels and face zones. In addition, the Hlotse drive required refrigeration equip- ment which provided chilled water to the tunnel. While the sec- ondary ventilation systems play an important role in gas and dust handling, the paper concentrates on the primary ventilation and cooling issues. The ventilation of these tunnels was an exacting exercise be- cause: • Rock temperatures and geothermal heat flow were high. • TBMs with relatively high power ratings were used. • Diesel locomotives were used. • Drives were relatively long. • High altitude meant a low air density. An important feature was the simulation and monitoring of the ventilation and heat flow components and the project was characterised by analysis, monitoring and ongoing tactical decision-making throughout the progress. The thermodynamics of the systems were complex because there were many interactive effects and analyses were carried out using special computer pro- grams. The monitoring confirmed the accuracy of the models, and in this manner it was possible to confidently ensure healthy and safe working conditions and still minimise costs. Local ambient climate conditions range from temperatures higher than 35 "C in summer to below -10 OC in winter. Based on available statistical data and the thermal storage/damping effects in the system, design summer ambient conditions were taken as 15 OC/25 "C wet-bulb/dry-bulb. The barometric pressure was 80 kPa and due to the altitude, the ambient air density was only 0,9 kg/m3. The local Authority specified a maximum in-tunnel wet- bulb temperature [at any point] of 32,O OC and a mean wet-bulb temperature [from all locations] of 27,5 OC maximum. The maxi- mum height of ground cover above the tunnel was 1 200 m and the maximum virgin rock temperature was 41 OC; see Figure 1. Diesel dilution criteria specified by the local Authority was a minimum of 0.1 m3/s per rated kW of diesel engine. Other requirements related to gases such as CO, CO2, NOx and CH4 [and the need for intrinsically safe equipment] but these are not of direct relevance to this paper. The actual average face advance was about 30 m/d with good days achieving 60 m/d and good months achieving 1 000 m [23 working days]. The original design tunnelling rate was 50 m/d. DESCRIPTION OF HLOTSE DRIVE VENTILATION AND COOLING SYSTEM The ventilation requirements in the tunnels were dictated by heat and diesel dilution needs. The best ventilation and cooling policy is generally a balance between using increased quantities of fresh air or refrigeration [or both]. In this particular scenario it turned out that, since the diesel emission criteria required large quantities of air, the refrigeration needs were modest. The drive was ventilated using a ducted, forced ventilation system from fans located at the portal. The maximum ventilation requirement was 51 m3/s when the drive was at 18.4 km. From a heat flow point of view, the worst scenario was a heat load of 3.5 MW when the drive was at 7 km. This was cooled by the ventilation air and a supply of chilled water to the tunnel. Refrigeration and chilled water system In the design phase, a detailed comparison was carried out between two general alternatives for providing refrigeration. First, was a system in which refrigeration sets and air coolers are installed on the TBM train; the refrigeration sets are cooled by condenser water piped to and from cooling towers at the portal. Second, was a system in which refrigeration water chillers are in- stalled at the portal and chilled water is piped into the tunnel. The detailed comparison indicated that the capital and running costs of the second system were at least 60 % lower than the in-tunnel plant. There were also many obvious practical benefits for favouring the portal system. The refrigeration plant supplied 23 11s of cold water at a temperature of 10 OC. After providing the cooling effect in the drive, the water returned to the portal where it was initially cooled in open-circuit evaporative pre-cooling towers, chilled in the refrigeration plant and then returned to the tunnel. The cold water flowed into the tunnel in an insulated supply pipe and returned in an uninsulated pipe; the water was simply circulated to the end of the pipe and returned. The cooling effect in the tunnel was achieved entirely through heat transfer from the pipe [long linear heat exchanger] and no air coils or other heat exchangers were required. The cooling requirements were satisfied by the heat transfer to the returns chilled water steel pipe [200 mm]. The pipes were eventually installed to a maximum distance of 10,8 km in what was considered a very practical and cost effective solution.

Jan 1, 1997

By Rolf Pollak, Jan Tegtrneier, Gerhard Keller, Horst Gerhardt, Wolfgang Müller, Dieter Tetzner

INTRODUCTION Under certain geological and mining conditions mining near the surface results in an intensive destruction of the overlying rock strata. If the system of mine openings remains filled with air after having completed all mining activities because the flooding of the mines depending on the morphology of the mining area does not exceed a certain level, this circumstance might result in dangers for the public. This is particularly true where uncontrollable air- ways occur within the destroyed rock or if former uranium mining areas are concerned. In these areas radon emanation resulting in undue radon concentration should be taken into account. This circumstance can be counteracted by suitable measures of ventilation. In the following the example of a uranium deposit in the Erzgebirge Mountains (Saxon Ore Mountains) is used to illustrate how statements about the ground flow through overlying rock can be made on the basis of systematic investigations. Finally comparisons with experience gained in coal mining yield orientation data which are used as input quantities in ventilation network calculations and in project preparation of mine fans. Last not least it is proved that endangering the population by radon can be excluded by installing an appropriate fan. INITIAL SITUATION From 1946 to 1989 the uranium deposit of Schlemal-Alberode was mined starting from superficial zones down to depths of 1,800 m. The total output amounted to 80,000 t of uranium, i.e. Schlemal Alberode is one of the largest uranium deposits in the world. The mine openings comprise a cavity of 40 million m3. Al- though thin seams (0.2 to 0.3 m) and steep to semi-steep veins (60 - 70 degrees on average) were extracted using the method of overhand cut-and-fill there were also deposit sections with larger thicknesses and poor ore zones, respectively, where especially after the war until about 1955 large openings were formed due to missing filling. Due to this uncontrolled, relatively "wild" mining considerable subsidence of the surface and fractures in the border zone of the subsidence area occurred. The town of Oberschlema was destroyed and an area of 0.14 km2 was damaged by mining involving a maximum subsidence of 6 m in the centre - called deformation area. The rehabilitation work for the deposit area of Schlemal Alberode involves the flooding of the openings up to a level of the Markus-Semmler-Sohle as long-term preventive measure. Fig. 1 shows the flooding level of October 1996. The level of the -540 m drift will probably be reached in June 1997. From that point of time a general rearrangement of the ventilation has to be made because the central return air ventilation via shaft 373 will no longer be possible. This shaft has no connection with the superficia1 system of mine openings. A new return air ventilation system based on shaft 382 as return air shaft will be designed (Fig. 1). The capacity of the future mine fan system will e.g. be determined by the flow conditions in the deformation area mentioned above. For this reason investigations were carried out in a test drift 13a on the level of the Markus-Semmler-Sohle in the stretching zone of the subsidence trough for several days (Fig. 2 and 3). INFLUENCE OF THE OPERATING CONDITIONS OF THE MAIN MINE FAN ON THE RADIATION SITUATION ON SURFACE The beginning of flooding in 1991 was also the beginning of the investigations into the options of future ventilation due to the reduction of the mining area. The first analyses of radon emanation made in May 1991 showed that only about 20 % of the total radon emanation result from systematically ventilated mine workings. This means that despite of existing dams the air streams through the majority of mine openings of the deposit. The analysis of the individual radon inflows illustrates the special problematic nature of the deposit involving all possible effects on the surface. The shutdown of the main mine fan over the long period between 2.10. and 6.10.1991 showed the essential influence of the operating conditions of the fans on the whole radiation situation on surface. Fig. 4 shows the comparison of the radon concentration level and its daughters during a 'normal" weekend with the level during long-term shutdown in the old 'Casino" of the papermill of Niederschlema. This papermill as well as the 'Casino" are situated in the outcropping zone of the deposit section of Niederschlema. The -60 m level was driven as the upper level in this area, from which the panels were driven to about 20 m below the surface. Also below the "Casino" there is such a panel in the vein 'Glirck", whose roof is directly situated below the building in a depth of 39 m. Directly next to it there is the raise 2 of the panel 'Namenlos" in a depth of 27 m. The graphics clearly shows the immediate increase in radon concentration as well as in the daughter concentration after the shutdown of the fans. The elimination of the artificially generated negative pressure results in the reversal of the convection direction due to the natural ventilating pressure. The interface structure allows the mine airto pass through aero- dynamic connections to the surface. Neglecting the meteorological parameters the concentrations of radon and its daughters in- creased by about 350 % as compared to a 'normal" weekend during longterm shutdown (i.e. in 4 days). Furthermore it could be noticed that a stationary state between the radon and its daughters has not been established in the period

Jan 1, 1997

By B. Elevli

Production scheduling of an open-pit mine is one of the most important steps in open-pit mining. The production schedule of a mine defines the distribution of cash flows and as a result directly influences the net present value (NPV) and overall economic feasibility of a project. Although the need to plan and operate a mine with an optimum production schedule has been obvious for some time, the methods used to obtain such a schedule have not been. The current practice for doing is to do it by hand on the basis of trial and error analysis. In an attempt to improve the way production scheduling is currently done, a new computer based production scheduling algorithm is developed. This paper describes the fundamentals of this newly developed heuristic scheduling algorithm and its application on a gold deposit. The results coming from a case study show that the schedules based on the new method provides improved cash flows compare to a manually done production schedules.

Jan 1, 1989

By Wallace Jarman

Introduction As with many gravity concentration processes, pneumatic concen¬tration traces its origin in antiquity. The use of winnowing to separate chaff from grain has long been known, and such procedures were undoubtedly practiced in ancient civilizations for the concentration of ores and the separation of slag from metal whenever specific gravity, size, and shape differences were favorable. Taggart142 has reported on the use of dry rockers and dry panning techniques for use with gold ores in arid regions such as Western Australia. The present day air table was developed in the last century for this purpose. Devices and Processes Classification. All classifiers make a separation on the basis of size, shape, and specific gravity. In addition to specific gravity and size, particle shape factor is often particularly important to pneumatic processes, and advantage is taken of the fact that both flat and fibrous particles settle at velocities substantially lower than that of their equiv¬alent spheres. Another factor of importance in this process is the bulk density of the material since substances such as exfoliated ver¬miculite or partially opened fibrils of asbestos have bulk densities significantly lower than those of the pure in situ mineral. Winnowing takes advantage of all of these physical factors to affect a separation starting with a closely sized feed, and vermiculite has been separated from rock by this process.142 When air classifiers are used to size minerals, the separation will often not be perfect because heavy particles will respond similarly to somewhat coarser light particles while flat, elongated, and low bulk density particles will act as would particles finer than the cut size. An example of a classification process that also results in concentration is the zig-zag classifier"' used to separate paper from glass, metal bottles, and cans during the recycle of municipal solid waste. Application of this process to minerals showing the proper specific gravity, size, and shape differ¬ences is obvious. An important device of this type is an air-aspirated screen used to remove asbestos from its associated rock during the dry processing of that mineral.144 The raw ore is crushed and screened and the asbes¬tos removed by aspiration from the surface of the screen by virtue of its shape and low bulk density. Hindered Settling. Based on the discussion in the section on "Hindered Settling Concentration and Jigging," the use of air devices could be anticipated in desert areas or where moisture may be deleteri¬ous to the product to be separated. Such processes have three major defects: (1) dust may be difficult to contain, (2) fines are difficult to process, and (3) the process is inherently less efficient than are wet processes as may be seen from Eqs. 3 and 4 in the section on "Hindered Settling Concentration and Jigging p. 4-47." Taggart142 has reported on the use of both dry panning and a dry rocker for gold separation in and regions. The dry panning makes use of winnowing, defection on the basis of specific gravity and shape during fall, and hindered settling. Prescreened placer material is poured from one pan to another as air blows across the falling stream. The process is repeated many times with hand picking to produce a rough concentrate. Magnetic minerals are then removed, and the residue, one grain deep in the pan, is air blown. In dry rocking the gravel is sized on a steeply inclined screen with the undersize fed to a riffle box with a porous bottom which is blown from underneath."' This device was the forerunner of the present-day air table. Air Tables and Jigs.145. 146 At least nine types of these devices have been developed, all but two of them for coal. While the devices used for coal may be characterized as pneumatic jigs, pneumatic tables, and pneumatic landers,145 only pneumatic jigs survive today. The same is true for ores.'" In these devices presized particles are fed to the separator which consists of an inclined vibrating conveyor with a porous surface through which air is carefully introduced to form a fluid bed. The lighter particles are lifted by the air out of uphill conveyance and float downhill, while the heavier particles in contact with the surface are conveyed uphill by the vibrator. Typically both transverse and longitudinal slopes are used. A plan view of a Triple S air table and the type of separation made is shown in Fig. 44.147 Because of its external appearance, the device has been called an air table, although it functions essentially as a jig. However, some of the asymetrical acceleration conveying function of a standard shak¬ing table is also performed by the eccentric vibrator incorporated into this device. Low pressure air is admitted below the vibrating table surface consisting of cloth (e.g., canvas), porous plastic, woven wire, or punched metal supported on a wire mesh or grid. As with a shaking table, the device produces a concentrate, middlings, and tailings. The middlings are recycled or are subsequently treated on another concentrating device. As with many wet separations, presizing by screening is necessary. The oft-repeated admonition of Professor Richards, "Separation without classification is damnation," is cer¬tainly true here. In this way a small heavy particle which might weigh the same as a large light particle and thus report to the same place on the air table has already been removed in the sizing step. Air tables are almost universally encased with a canopy to remove dust by aspiration. Some concentration may also be achieved in this step when favorable specific gravity, size, and shape differences are present. Applications Ore Separations. Air tables were originally developed for ore separations, and they find application in and regions or where water is deleterious or inconvenient to use or remove (e.g., small tonnage materials already dried). They have been used for asbestos, bauxite, calcite, cassiterite, columbite-tentalite, diatomaceous earth, fluorite, gilsonite, graphite, kyanite, manganese minerals, mica, managite, perlite, pyrite, pyrrhotite, vermiculite, and uraninite ores. More recently, air tables have found favor in separating a wide variety of secondary materials such as abrasive grains, bone char, catalysts, fiberglass, scrap glass, scrap wire from its insulation, prills from slag, dross and coke from metal or from each other, metal from crushed crucibles, lead from plastic in old batteries, and cubic particles from flat ones. The devices also finds considerable application in the preparation of food¬stuffs. As stated before close sizing before separation is desirable, and if the specific gravity difference is slight, very tight

Jan 1, 1985