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By R. K. Singhal

In open pit mining every effort is made to maintain as steep a pit slope as possible. Pi t slope angles are influenced by numerous factors some natural that are amenable to limited or no control at all and other mining technology related where a greater degree of influence can be exerted by the engineer. One such area where control is possible is the method of drilling and blasting in the pit. This paper discusses the influence of blasting techniques on the economics of pit slopes and demonstrates how balanced blasting can be used with advantage to pit wall control and improve fragmentation (thus reducing material handling costs). Microprocessor based instrumentation to monitor drill and blast performance has also been described.

Jan 1, 1988

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 Richard E. Ray

INTRODUCTION Main mine fan installations at underground coal mines must be designed to comply with the requirements outlined in 30 CFR (Code of Federal Regulations) Part 75. However, very little emphasis is typically placed on the aerodynamic optimization of shaft collar designs, ductwork configurations, and the selection of a fan isolation, or "closing door," system. Field testing has shown that poorly designed systems may produce large pressure losses between the shaft and fan inlet, as well as air turbulence problems which adversely impact fan efficiency and sometimes fan structural integrity. This paper analyzes each component of a surface exhaust fan system for an underground coal mine: • The shaft collar arrangement, • The 45 or 90-degree turning duct that isolates the fan motor from the return airstream, • Alternative ductwork arrangements for parallel fan installations, • The self-closing doors or dampers that are required for multiple fans, and • Horizontal and vertical evas6 arrangements on the discharge side of the fan. While discussion is limited to axial fan installations at shaft mines, most of the recommendations would also apply to centrifugal fan installations and to main mine fans for drift mines. FEDERAL REGULATIONS The mandatory safety standards for underground coal mines that apply to the installation of main mine fans are listed in 30 CFR Part 75.310. Stipulations which impact the design of the fan sys- tem arrangement are contained in paragraphs (a)(l), (a)(5), (a)(6),(d), and (1) of 75.31 0: • Each main mine fan shall be installed on the surface in an incombustible housing; • Each main mine fan shall be protected by weak walls or explosion doors in direct line with possible explosive forces; • Explosion doors or weak walls shall have a cross-sectional area at least equal to the area of the return air shaft or drift opening through which the pressure from an underground explosion would be relieved; • Each main mine fan shall be offset by at least 4.57 m (15 feet) from the nearest side of the mine or shaft opening; • Automatic closing doors shall be provided to prevent possible air reversals through the fans in mines ventilated by multiple main mine fans. The installation of main mine fans underground is prohibited by 30 CFR Part 75.310, (a)(l). Providing explosion doors or weak walls in line with a possible underground explosion and offsetting the fan 4.57 m (1 5 feet) from the edge of the shaft preclude the installation of the fan directly over the shaft. In addition, 30 CFR Part 75.507-1 specifies that all electrical equipment used in return air courses be permissible, which requires locating the fan motor outside of the airstream exhausted through the fan. To comply with the Part 75.310 regulations, explosion doors are typically located directly over the shaft as part of the shaft cover or 'bonnet." A sufficiently long section of ductwork placed between the shaft enclosure and fan, equipped with a 45 or 90- degree turn, enables the fan to be offset at least 4.57 m (1 5 feet) from the shaft and the motor to be isolated from the airstream. The need for the turn in the ductwork can be eliminated by driving the fan from the discharge side, instead of from the inlet side, and providing for a 90-degree bend in the evas6. For multiple main mine fan systems, self-closing doors can be located on the inlet or discharge side of the fan. FAN SYSTEM COMPONENTS A poorly designed exhaust fan arrangement will produce turbulent conditions at the fan inlet and significant pressure losses between the shaft and fan. Field tests have shown that pressure losses through inlet ductwork at some installations may constitute as much as 30 percent of the total mine head. Unfavorable flow conditions may cause a change in the operating point on the fan performance curve, resulting in reduced fan efficiency and pres- sure generation capability. In some instances, poor airflow distribution through the fan may create mechanical stresses that lead to fan blade or fan housing fatigue failures. The objective in designing an air duct arrangement for a fan system is to provide a low cost duct that will move the air from the shaft to the fan in a uniform air distribution and with minimum energy requirements. The optimum solution is a compromise between technical and economic factors. Aerodynamic design considerations for the shaft collar and cover, fan inlet ductwork for both single and parallel fans, self-closing doors, and evas6 are discussed in the sections below. Shaft Collar/Cover To prevent excessive turbulence and non-uniform air distribution through the fan, a smooth transition for the change in direction of airflow at the top of the shaft must be provided. However, the concrete lining of the shaft is abruptly terminated at the surface at many coal mine installations (see Figure I), creating a sharp 90-degree bend. The airflow exiting the shaft becomes heavily concentrated in the upper portion of the horizontal fan in-

Jan 1, 1997

By William Lyman

GENERAL DESCRIPTION Shrinkage or shrinkage stoping refers to any mining method in which broken ore is temporarily retained in the stope to provide a working platform and/or to offer temporary support to the stope walls during active mining. Since ore "swells" when broken, it is necessary to shrink the muck pile a corresponding amount by draw¬ing some of the broken ore out as the stope is advanced-hence the name. Broken ore retained during stoping is drawn out after the stope has reached its limits. The stope may be left empty or may be filled with waste contemporaneous with, or subsequent to, the final draw. Traditionally the method implies conventional overhand stoping methods with miners working between the muck pile and the stope back, in a space which advances updip with mining and is maintained by balanc¬ing "swell" with "shrink." The shrinkage classification is also applicable to so-called "semishrinkage" methods in open pillar-supported stopes where broken ore is temporarily retained as a working platform but offers no wall support; and to various blasthole shrinkage methods which utilize broken ore temporarily retained in the stope for wall support, but which do not require miners to work from muck pile in the stope. The method is generally applied to steeply dipping veins of strong ore between strong walls. APPLICATION Geometry The geometry of a shrinkable vein is described in terms of dip, width, and regularity along dip. Overall strike and dip dimensions and irregularities along the strike generally impose no restrictions on the method. Dip is ideally 1.2 to 1.5 rad (70 to 90°). As dip falls below 1.2 rad (70°), the shrinkage draw begins to strongly favor the hanging wall side, thus leaving a poor working platform for conventional overhand work. This is particularly true in relatively wide stopes. The sup¬port afforded to the hanging wall also diminishes with decreasing dip, reaching nil as the dip approaches the repose angle of broken ore. Dips below 0.78 to 0.87 rad (45 to 50°) are not generally shrinkable except by open stope "seinishrinkage" methods. Minimum mining width is fixed by working space requirements in the stope-generally about 1 m. Shrink¬age in narrower veins requires that waste rock from one or both walls be broken with the ore and the attendant dilution accepted to achieve the minimum width. Nar¬row stopes are less suitable, encouraging hang-ups and bridging of broken ore, with the attendant problems of erratic draw and incomplete recovery of broken ore. Maximum practical width may be 3 m or less to over 30 m, depending upon the competency of the ore and its ability to stand unsupported across the stope back. This is a vital safety consideration in conventional over¬hand stopes, but is much less of a factor in blasthole shrinkage methods. Very wide veins and massive ore bodies have been mined by transverse vertical shrinkage panels separated by transverse vertical pillars which are either abandoned or recovered later by other methods. Regularity along the dip is a prerequisite of shrink¬age as there must be no serious obstruction to the flow of broken ore downward through the stope to the sill level. Gentle rolls along the dip are acceptable if the local footwall dip everywhere exceeds 0.78 to 0.87 rad (45 to 50°). Off-dip hanging wall and/or footwall splits can generally be mined selectively from a conventional shrink stope as they are encountered without ad¬versely affecting subsequent continuation of shrinkage mining updip on the main vein. Vertical offsets or major rolls along the dip which cannot be "smoothed over" generally require that a sublevel be established with new draw control development. Blasthole shrinkage methods are much less flexible (and thus less selective) in their ability to accommodate any of these irregularities. Ground Conditions The wall rock must be strong enough to stand with the minimal support afforded by the dynamic mass of broken ore in the stope. During active mining, local sloughing from the walls is restrained, but the broken ore affords little, if any, useful resistance to closure of the stope walls. Such squeezing, if present, may bind up the stope and cause the loss of much ore. Pillars left between and/or within stopes are effective in preventing closure but reduce overall recovery. Walls may be re¬inforced by bolting after each stope cut in conventional shrinkage but not in blasthole shrinkage. Ore in place must be strong enough to stand with no natural support across the stope width, although tem¬porary artificial support or reinforcement may be used locally in conventional stopes. Some spalling or sloughing is permissible in blasthole shrinkage as men are never present in the stope. Physical and/or mineralogical characteristics of the broken ore may impose restrictions on stope design and/or operational plan¬ning, and may even preclude the use of shrinkage al¬together. Examples include: ores which, when broken, are cohesive or which tend to pack or cement together under the influence of ground water, wall pressure, and/ or chemical reaction. Such conditions precipitate er¬ratic draw during mining and often result in difficult and/or incomplete final draw; pyritic ores which oxidize very rapidly in the stopes and may generate heat, imposing a fire hazard by spontaneous combustion; sulfide ores which oxidize sufficiently in the stopes to adversely affect mill recovery by flotation; and ores (es¬pecially those containing uranium minerals) which ex¬ude radon gas and thereby impose ventilation constraints on stope design. In most cases these problems can be minimized by limiting the size of stopes, by minimizing the duration of mining activity in each stope, and by promptly drawing each stope empty following comple¬tion of mining.

Jan 1, 1982

By Norbert Patzelt, Johann Knecht

INTRODUCTION Grinding mills have been used in the minerals processing industry for over 100 years. Their dimensions have grown continuously during this time. Besides increasing throughput rates of grinding plants due to the depletion of high grade ores, the lower specific in- vestment costs, as well as reduced operating and maintenance requirements are major reasons for this trend. When selecting new plant equipment one must consider that design principles which have proven their reliability on sizes of today's equipment do not automatically warrant a successful operation on the ever larger size of equipment. Modern calculation methods as for instance the Finite Element method already contribute considerably to the safe design of the huge equipment being built today and are a standard tool of the design engineers. More recently, modern computer programs are also being used in order to size the equipment to meet the process requirements. Today, two design principles are on the market - one which supports the weight of such a unit on trunnion bearings through cast conical endwalls and one which is supported through slipper pad bearings arranged at the circumference of the mill shell (Fig.1). The reason for the development of this alternative grinding mill design can be found in the past. During the sixties and seventies the growing sizes of ball mills with high LID ratios caused many mill failures due to cracked endwalls. The accuracy of the calculation methods as well as the quality standards for castings were not developed to a degree required for such kind of heavy equipment. One way to overcome these problems was the increase of the manufacturing quality standards as well as the introduction of the finite element method based on the analysis of the experience available. The biggest grinding mills being built today are large size SAG mills with cast conical endwalls and trunnion bearings (Fig.2). This is due to the fact that mill manufacturers who had come from the conventional ball mill design adopted these principles as well to their SAG mills. These grinding mills perform well without special concern to the operators. Other manufacturers overcame the problems as mentioned above by eliminating completely the heavy castings and trunnion bearings and the problems associated to it (Fig.1). This design was originally applied to ball mills for the mining and other industries. Due to the success of these shell supported ball mills, this design principle was also applied to SAG mills(Fig.3). Despite of the fact that the majority of today's grinding mills are built to the conventional design it is also interesting to have a look at this alternative. Principles which have proven their reliability on sizes of today's equipment do not automatically warrant a successful operation on the ever larger equipment if bigger mill sizes are realized only based on the pantograph principle. With growing grinding mill sizes, the mass and volume flows through the equipment increases rapidly. Thus it is very important not only to concentrate on the safe design of the structural components of the equipment but as well on the process requirements. The influence of the design on important process parameters of dry and wet grinding plants are discussed thereafter. It shall be shown how modern computer programs can assist in the optimization of the design of components in order to fulfil the operational requirements of such large size equipment. PROCESS REQUIREMENTS OF LARGE SIZE GRINDING MILLS Dry Grinding Mills The world's biggest ball mill is a dry grinding ball mill having a diameter of 6.2m and an overall length of 25,5m with a drive power of 11,200 KW or 15,000HP. This grinding mill dries and grinds gold ore at a rate of 500 tons per hour at a moisture content of up to 9,5%. As shown in Fig.4 this mill was built as a shell supported unit. In fact only this design principle allowed to meet the process requirement. This mill could hardly be built with cast conical endwalls due to the constraints of the trunnion bearings limiting the mill inlet. The following case shows how modern computer programs can help to meet the design criteria of the air system of large size dry grinding plants. For dry grinding plants, the gas flow through the SAG mill has to match the drying, as well as the material transportation require-

Jan 1, 1998

By G. D. Bates

INTRODUCTION The Magmont mine is a joint venture of Cominco American Inc. (operator) and Dresser Minerals, Inc. The mine-mill operation is located approximately 160 km (100 miles) southwest of St. Louis, MO, in what is commonly referred to as the "Viburnum Trend.” The Magmont mine is designed for a production rate of 3810.2 t/d (4200 stpd) on a 5-day week, three shifts per day basis. Initial production began in 1968. The mine is open stope, room-and-pillar, and essentially horizontal along the trend of the ore body. Briefly, the main geological structure can be described as a brecciated graben bounded by reverse faults. The ore body in cross section is shaped like a bell curve with some lateral extension at the lower part. Presently outlined ore is 609.6 to 762 m (2000 to 2500 ft) in width and 2133.6 m (7000 ft) in length. The ore varies in thickness from 4.87 m (16 ft) on the fringes to an average of 27 m (90 ft) in the high ore areas bounded by the reverse faults. Lead is the primary metal with zinc and copper secondary. MINE DESIGN The basic design of open stope, room-and-pillar mines has been described by several writers and need not be repeated here. (Anon., 1970; Bullock, 1973; Casteel, 1972; Christiansen et a]., 1970; and Lane, 1964) This discussion covers the mining sequence as applied to the particular problems at the Magmont mine, the use of equipment, and deployment of the work force. In the upper portion of the Magmont ore body is a layer locally called the False Davis shale. This layer lies below the true Davis shale, is normally interbedded with dolomite, is of varying thickness, and if mineralized, is included in the top pass of the mining sequence. In thick ore areas this layer will be 2.13 to 2.43 m (7 to 8 ft) in thickness and will occur in the upper portion of the pillars. Due to its incompetency the presence of this False Davis layer is of primary concern in mine planning and operation. Mining areas are divided into three basic groups by ore thickness. First are areas of ore up to 6.09 m (20 ft) in thickness. These areas are below the False Davis shale and are mined single pass with drill jumbo. Second are those areas up 13.71 to 15.24 m (45 to 50 ft) in height. The first 4.87-111 (16-ft) Pass is taken at the top of the ore and the back and pillars secured. Benching the lower portion(s) in 4.57 to 4.87-m (15 to 16-ft) passes is then done with either a drill jumbo drilling horizontally or a crawler drill drilling vertically. Normally these areas are below the Table 1. Productivities per Manshift False Davis shale. These areas may also be mined by back slashing, or overhand benching, where the first 4.87-m (16-ft) pass is taken at the base of the ore and successive 4.87- m (16-ft) passes are taken upward. A minimum amount of back slashing is done at Magmont since it requires repetition of roof control on each pass and roof control is the single largest stoping cost at Magmont. Ore left to provide a working platform oxidizes and is coated by oil spills thus reducing metallurgical recoveries. The third mining area is over 15.24 m (50 ft) in height UP to a maximum of 40.23 m (132 ft) and will encompass the False Davis shale. These areas are mined by first driving +15% inclines to the top of the ore body. The top pass is mined and the back is bolted and roof mats installed as a matter of standard practice to minimize roof problems as mining progresses downward. Once the back and pillars on the top pass are secured, benching begins on successive passes with either the drill jumbo or crawler drill. Pillars on all successive passes below the top pass are secured as necessary. While benching progresses below the top pass, the pass at the base of the ore body is mined leaving a sill of 4.57 to 7.62 m (15 to 25 ft) in thickness to be removed with the crawler drill in a retreating manner. Rooms are mined on a 1.57 rad (90") grid pattern to insure alignment of pillars where multiple passes are taken. Pillars are designed on a 17.98-m (59-ft) spacing with rooms up to 10.66 m (35 ft) in width. Heading widths are wide enough for the mobile equipment to turn without additional allowance for curves. The result is a flexible layout which provides a maximum number of headings available for high extraction rates and grade control. PRODUCTION Incentive Bonus Incentive bonuses play an important part in the mine production at Magmont. Production crews are trained to perform only one of the mining functions of drilling, blasting, mucking. or roof bolting. This specialization, or functionalization, is augmented by development to open all possible stoping areas as early as possible in the life of the mine. This insures that each crew will have enough headings to perform its specialty. The incentive bonuses increase exponentially as output increases. The lucrative incentive bonus coupled with the specialization of the production crews and proper mine development have combined to give the high productivities shown in Table 1. Development crews perform all mining functions in their area. The incentive bonus is paid on a per foot basis, Crews on different shifts working the same heading share equally in the bonus proportional to their contract hours worked.

Jan 1, 1982

By James R. Arnold

Over the past few years several commercial wetting agents and drying aids have been tested in the lab and in the plant at Gold Fields Mining's operations at Mesquite and Ortiz. The major goal was to increase recovery with these surfactants by improving the permeability of the pad and increasing the rate of penetration through the rock.

Jan 1, 1988

By Roy M. Fleming, Christine B. New

The federal standard for limiting exposures to miners from radon daughters was reduced from 12 working level months (WLM) per year to 4 WLM per year in 1971. However, even at that time some researchers were concerned that the new limit would eventually be shown to result in an excess risk for lung cancer mortality in miners. The National Institute for Occupational Safety and Health (NIOSH) is now engaged in a comprehensive review of this topic. As the lead agency in this project, NIOSH has developed a work plan and established a work group to implement this plan. The procedure and specific considerations are outlined in the work plan for developing a comprehensive ionizing radiation standard recommendation for all miners, underground and surface. Such a recommendation will not only include estimates of health risks at various levels of exposure, but also appropriate recommendations for medical monitoring, sampling and analytical methods, sampling strategies, posting, engineering controls, personal protective equipment and recordkeeping. The work group has twenty-four members. Eleven members are NIOSH personnel representing six divisions of the Institute. The remaining thirteen members represent other federal agencies, specifically the Bureau of Mines of the Department of the Interior, the Mine Safety and Health Administration and the Occupational Safety and Health Administration of the Department of Labor, the Office of Radiation Protection of the Environmental Protection Agency, and the Bureau of Radiological Health of the Food and Drug Administration, Department of Health and Human Services. Several factors contributed to the decision to initiate this project. First, the efficacy of the current standard was considered. An initial study group was formed in the spring of 1980 by NIOSH to identify and evaluate articles that contained information on lung cancer mortality risks at and below the present permissible exposure limit. The conclusion drawn from their evaluation was that a two-fold excess risk of lung cancer mortality at and below 120 cumulative working level months (CWLM) of exposure to radon daughters is evident. This composite indication from selected studies was of sufficient magnitude to justify further evaluation. The study group, however, recognized that other studies and information must also be considered in a quantitative risk assessment which would form the basis for recommending an acceptable and feasible permissible exposure limit. A second factor in the decision to pursue further evaluation was the gaps in the current standard that had previously been identified by the Mine Safety and Health Administration (MSHA). These included lack of medical monitoring of underground miners and absence of regulations for surface miners. The seriousness of the health hazard relative to other hazards in mining was also considered. Along with exposures to silica and asbestos fibers, radiation was judged to be one of the major health hazards in mining, These combined factors constituted the justification to develop criteria and recommendations for improved mandatory health standards. The development process begins with a survey and review of the available world-wide information on the topic, including data and information developed by NIOSH. The end product of this review is to be a document that will contain an evaluation of the collected information and support for any recommendations that are made. To develop this document, the present work group has been divided into five task groups with the following emphases: Health Effects, Medical Aspects, Monitoring, Environmental Exposures, and Engineering Controls and Work Practices. The Health Effects task group is to evaluate the evidence from epidemiologic and animal studies of adverse health effects associated with all forms of ionizing radiation encountered in mining and milling operations. The "weight-of-evidence" of the results of all relevant and useful studies will be summarized, with the reasons for emphasizing the cited studies. The critical cells or tissues and the factors that should be considered in estimating the dose to these areas from various types of radiation will be identified. An evaluation will be made of the possible impacts of smoking and exposure to diesel exhaust on the determination of health effects related to radiation. The implications of biologically redundant dose in terms of the time between tumor initiation and death will also be analyzed. The Medical Aspects task group is to review the generally-accepted and the state-of-the-art medical technology for the detection of adverse health effects from ionizing radiation exposure. An evaluation will be made of the accuracy of urine and fecal analysis, wholebody counting, chromosome analysis and nose blows, as well as their usefulness for early detection of adverse health effects. Early detection is of little utility to the affected individual unless subsequent medical care can improve the prognosis. Recommendations for screening tests will be made after carefully considering the accuracy of the diagnostic procedures and the usefulness of early detection. Required recordkeeping and transfer rights will also be addressed. The Monitoring task group is to consider the state-of-the-art technology for the monitoring of occupational radiation exposures. Instrumentation, sampling strategies and analytical procedures will be reviewed for both personal and area sampling. The implications of the associated levels of confidence for non-compliance decisions will be evaluated. The discussion will also include an evaluation of the feasibility of replacing present monitoring systems with recent technology and the impact that a possible lower permissible exposure limit would have on monitoring requirements. The Environmental Exposure task group is to investigate the field procedures and mathematical methodologies which have been used to quantitate exposure levels to the various kinds of ionizing radiation. The magnitude and direction of the possible biases in past exposure assessments will be estimated. Specific attention will be given to the controversies concerning the quantification of the biological dose equivalent and the feasibility of recommending standards for mixed radiation exposures which use "rem" as the unit of measurement. The Engineering Controls and Work Practices task group will analyze the advantages and limitations of

Jan 1, 1981

By R. S. Dennen

A ground penetrating synthetic pulse radar system has been used by the Bureau of Mines to probe into a coal pillar to locate hidden structures or anomalies. Two types of velocity tomographic reconstruction methods, direct matrix inversion and iterative, were used to obtain a velocity image of the interior of the pillar. The reconstructed image showed the existence and location of a relatively low velocity material within the pillar. Core samples were taken from the pillar in a region surrounding the low velocity material. Analysis of the cores showed the presence of a clay vein coinciding with the low velocity material imaged in the tomogram.

Jan 1, 1989

By K. R. McGinnis

A major cost to a mining company is funding of an exploration program. It is a necessary cost to insure the company's future. But if the exploration effort does not discover an economic are body, there is no return on investment and the days of the company are numbered. However, if you are aware of what exploration geologists are saying, it places the exploration process in a different light. Basically they are saying that ore bodies exist that contain 250,000 tons or less of gold and silver bearing ore that is cyanide amenable utilizing heap leach technology.

Jan 1, 1988

By R. R. Rollins

Conventional methods of controlling abandoned mine fires include: loading out, inundation, fire barriers, flushing, and surface sealing. This paper describes a proposed fire-control technique utilizing a new explosive fragmentation method. An impervious barrier is created by using explosive energy and strategic emplacement in boreholes drilled from the surface. Blast round design, average fragment size, and impermeability of the fragmentation barrier are discussed. This barrier, created by the blasted floor rock, could also be used to fill in the open space in abandoned mines for subsidence prevention or control.

Jan 1, 1989

By A. Mukherjee

This paper presents an analytical technique for prediction of the damaged rock zone (DRZ) around a circular drift subjected to an anisotropic stress field (analytical solution exists only for the case of hydrostatic stress field). The method can readily be adapted to micro or personal computers. Stress concentrations around the drift are calculated at the centers of small rock elements, using the analytical solutions for a hole in an infinite plate. Potential failure of a rock element is assumed to be governed by the Coulomb failure criterion. When the failure condition is exceeded for an element, the excess shear stress is transferred to the neighboring elements. Progressive failure is allowed to occur until convergence is obtained. The final configuration of the damaged rock zone so obtained can be used to assess rock support requirements, as well as modified modulus and permeability of the DRZ.

Jan 1, 1988

The Witwatersrand reefs provide excellent examples of chemical and structural control on gold distribution. All mined gold appears to coma from a carbon -gold association and/or Fe-S-Au association, and also from immediately above or below unconformity surfaces. A total raliance on the current syngenetic model for Witwatersrand gold does not appear to be justified from the accumulated data and may be responsible for missed opportunities during exploration: consideration of other models is suggested especially where they better fit the data. The stratabound natura of Witwatersrand gold 1& critical to mining, but is emphasized in syngenetic models to the virtual exclusion of many other important observations. By analogy with other deposit types, stratabound character is probably a poor guide to genesis.

Jan 1, 1988

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

The STG integrated method of grouting can be divided into the several phases as outlined in preceding chapters. For convenience in describing technical procedures, they can be grouped as: 1. Investigations in non-geotechnical, exploration bore¬holes. 2. Drilling the holes necessary for geotechnical investiga¬tions and for grouting. 3. Conducting appropriate hydrogeologic tests in those holes and consequent calculations. 4. The preparation of a clay mortar. 5. The preparation of a clay-cement grout with additives. 6. The injection of the grout. 7. Checking the quality of the grout curtain. Items 2, 4, 5 and 6 are discussed below. These activities contain the procedural aspects of grouting. 7.1 DRILLING GROUT HOLES Grout holes belong to a very select class of drillholes. The necessary drilling equipment must enable the operator to drill the holes in inclined, twisting directions. In addi¬tion, the equipment must permit drilling under structurally and hydrogeologically complex conditions. The equipment must permit the operator to conduct specific testing activi¬ties in the drillholes. These activities include hydrodynamic analyses, flowmetric analysis, and the installation and re¬moval of deflectors, packers, grouting plugs, casings, lin¬ers, pumping facilities and other work. As explained in Chapters 5 and 6, the grouting of sat¬urated rock is conducted both through holes drilled from the surface and through holes drilled from the face of a shaft, drift, or tunnel. The drilling of grout holes from the surface can be carried out by an aggregate of equipment that can be the same equipment used for drilling exploration boreholes. Depending on the projected depth of the holes, the STG ZIF- 1200MR and ZIF-650 drilling rigs (or modifications of them) are used. The SKB-4, SKB-5 and SKB-7 high-output drilling rigs (workover rigs) have been used in recent years. The STG BMP-24 drilling rigs are used for carrying out the descent-lifting operations for drilling holes from the sur¬face. In a number of cases, the UKB-500C, URB-3AM and URB-2A power-fed drilling equipment is used for drilling holes from the surface. It should be noted that the drilling of grout holes usually requires an electric drive assembly. Drilling with diesel drive assemblies would be used only to preclude the possi¬bility of losing electric power during drilling operations. Turbine drilling would be used only for appropriate techni¬cal-economic reasons. The selection of the grout hole design is determined by the hydrogeological and structural geological conditions at the site as interpreted in Chapters 2 and 3. The design of grout holes must be as simple as possible in order to min¬imize costs. Hole design guidelines can be adopted as a basis for this purpose, unless site-specific conditions require otherwise. The mouth of the drillhole must be outfitted with a guide-pipe having a length of at least 2 m and an outside diameter from 219 to 234 mm. The upper part of the hole must be attached by a jig consisting of borehole casing that has a diameter ranging from 108 to 146 mm. The length of the jig h is determined by the equation [ ] where k = 1.1 to 1.2 is the load factor; P,, is the injection pressure of the grout into the mouth of the drillhole; D is the external diameter of the jig pipe; m = 0.6 to 0.7 is the work condition factor; T[ ].1 MPa is the bonding value of the rock cemented to the jig. The grout hole is drilled from the jig shoe to the de¬signed depth using a rock-crushing bit with a diameter of 93 mm or more. However in complex hydrogeological condi¬tions when the shaft intersects unstable rock, other hole designs can be used. The diameter of grout holes must permit an aggregate of investigations to be conducted in them using down-hole borehole geophysical and flowmetric logging instruments. It is necessary to cement the casing strings reliably in the grout holes, thereby permitting the injection of the grout under high pressure through the pres¬sure tight mouth of the borehole. Technological details for drilling grout holes such as the rotational rate of the drill bit, the axial load on the rock¬crushing bit, and the drilling fluid flow rate are optimized for the specific rock-hydrogeological conditions largely by experience. The hole diameter and type of rock crushing bit are important variables in these details of drilling. It is advisable to drill holes using water as the drilling fluid. However in those cases where the hole walls are unstable or under very high ground water pressure, the use of a well¬ engineered drilling mud should not be precluded. Unfortu¬nately, drilling mud can influence test results. When steeply dipping fractures with inclination angles of more than 60 to 70° are encountered, it is necessary to drill guided, inclined grout holes. This procedure permits the maximum number of fractures in each hole to be stu-

Jan 1, 1993

By George B. Rice

In a recent speech in Pittsburgh, Dr. George Keyworth, the President's Science Advisor, made a statement which I believe deserves our very careful consideration. Dr. Keyworth said that there is no energy crisis. The crisis, he explained, is simply that people refuse to accept the solution. The solution which Dr. Keyworth has in mind is increased utilization of our abundant supplies of solid fuels and, in particular, uranium. I share his view concerning the solution to our energy needs. The use of uranium fuel is a safe, clean, and dependable means to generate our electric power. It is time that we addressed the real energy crisis: the refusal to accept the nuclear solution. The reason for the refusal is not difficult to find. It is nihilistic thinking about risk. Under this thinking, we assume the worst possible case and act accordingly, simply because we cannot prove to a total certainty that nuclear energy is perfectly safe. If this absolutist approach were generally applied throughout our society, there is no doubt all of us would soon be sitting around our campfires fearfully holding the wild animals at bay with our trusty spears. Today I am here to enlist your support in reversing the regulatory trend that threatens the very extistence of the nuclear power industry. As distinguished scientists, engineers and businessmen, you can use your influence to help bring rational regulation to the industry. Our industry supports strong safety and environmental protection programs. We understand the need for and do not object to reasonable regulation. Many anti-pollution measures can be practical to implement, cost effective and highly successful in minimizing environmental impacts. However, it is a fact of life that in the field of health and safety regulation, the law of diminishing returns operates with a vengeance. Absolute or near-absolute safety is impossible and any attempt to achieve it is intolerably costly. Fixation on absolute safety is particularly acute in the regulation of the nuclear power industry. Government Agencies, overly anxious to allay the irrational fears of those opposed to nuclear power, are literally regulating the industry to death - exactly the result sought by the anti-nuclear groups. Dr. Robert L. DuPont wrote in a recent issue of [Business Week]: "The nuclear power industry has been virtually stopped in the U.S. [because of fear]. This is true despite the fact that for more than 20 years the commercial nuclear industry has operated under unprecedented public health scrutiny and that to date there have been no radiation-related injuries, let alone deaths, suffered by any member of the public."1 I believe a useful way to convey the nature of the problem faced by the nuclear industry is to review an example of [unreasonable] regulation. While the example relates to our domestic industry, I am certain there are similar situations in other countries. For the example I will use the Nuclear Regulatory Commission's recently issued regulations governing the stabilization of uranium mill tailings.2 These regulations, known as the Uranium Mill Licensing Requirements, specify, among other things, that radon emanation from uranium mill tailings be limited to no more than 2 pCi/m2-sec. First, one must understand that this standard will have virtually no impact on the total amount of radon to which the public is exposed. Radon emitted from even completely unstabilized tailings piles is a tiny fraction--much less than 1%--of the amount of radon released from natural soils in the United States.3 In fact, it is far outweighed by natural variations in the background flux. For example, changes in the level of the Great Salt Lake in recent years have had [eight times] as much effect on the amount of radon released into the Salt Lake City regional air than the annual release from the Vitro Mill tailings pile located in that city.4 Nevertheless, NRC claims that the standard is required to protect the public. The Commission admits, however, that there are no studies which establish that exposure to radon at the low levels associated with uranium mill tailings will result in any adverse health effects.5 In the absence of actual evidence, the Commission assumes that some such effects will occur on the basis of the linear, non-threshold model.6 Employing this model, NRC calculates that the maximum hypothetical risk for the average member of the population is only about 1 in 70,000,000 from the radon that would be emitted from [three times] the number of mills now in existence, even if the tailings produced through the year 2000 are left unstabilized.7 NRC has elsewhere explained that this level of risk would be equivalent to the risk posed by "a few puffs on a cigarette, a few sips of wine, driving the family car about 6 blocks, flying about 2 miles, canoeing for 3 seconds, or being a man age 60 for 11 seconds." This level of risk is [de minimis] in comparison to other risks commonly and readily incurred in our society.9 Moreover, even this remote risk is overstated. A group of prominent health physicists, including experts from the Department of Energy, The Environmental Protection Agency, Britain, Canada and Germany recently published a study indicating that the risk to the public per unit exposure to radon can be no greater than one-third that suggested by the Commission, and [may in fact be zero].l0 Regulators routinely rationalize the need for their regulations. For example, NRC attempts to justify the radon flux standard because it is necessary to reduce the risk to someone who builds a house on top of a tailings pile. This possibility, however, is totally unrealistic because the Mill Tailings Act requires that stabilized tailings be transferred to

Jan 1, 1981

By A. R. Przybylek

A fluid bed dryer has been designed for a nonconventional application to upgrade the value of fine anthracite culm and silt recovered from culm banks and anthracite tailings pond beds. This material can be seen as large spoil piles scattered throughout the anthracite coal regions of Eastern Pennsylvania. Engineering of the system commenced in December of 1986 and at this writing start-up is scheduled for late October or early November of 1987. This paper will focus on generalities concerned With system design and operation, and qualify the differences from conventional fluid bed coal drying systems.

Jan 1, 1988

By R. G. Robins

The dissolution of metallic copper powder in aqueous solutions of cupric sulfate with ethylene under low pressure was investigated. The reaction stoichiometry obtained was:

Jan 1, 1988

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

2.1 INTRODUCTION When total extraction of an opening of sufficient size is reached in a horizontal coal seam, the roof strata in the overburden deform continuously to reach a new equilibrium condition. The severity of deformation decreases upward toward the surface. As the downward saggings of the strata propagate and reach the surface, there will be a depression zone on the surface directly above, but extending beyond the edges of the underground opening. This is the surface subsidence basin or surface subsidence trough. The surface subsidence basin is circular in plan view, if the coal seam is horizontal and the mined-out opening is square in shape. But it is rectangular with rounded corners or elliptical if the coal seam is horizontal and the mined-out opening is a long- and thin rectangle or a short-rectangular, respectively (Fig. 2.1). Most underground openings (e.g., longwall panel) assume rectangular shape when total extraction has been completed. Theoretically the edges of the subsidence basin are the points of zero subsidence. But it is difficult to exactly locate the points of zero subsidence. Therefore in practice the points with vertical subsidence of 0.4 in. (10 mm) are used. The final subsidence basin is that which exists long after the mining has been completed, because its magnitude and shape are quite different from the dynamic subsidence basin formed while the face is moving. 2.2 CHARACTERISTICS AND TYPES OF DEFORMATION IN THE FINAL SUBSIDENCE BASIN For a horizontal coal seam, every point in the subsidence basin moves toward the center of the basin. Subsidence is maximum at the center of the basin. Any cross-section that passes through the point of maximum subsidence and either parallel to AB or CD line (Fig. 2.1) is a major cross-section along which principal directions of surface movements occur. However among those infinite numbers of major cross-sections, two specific ones are of special significance, not only because the magnitudes of surface movements are the largest, but also because they are the most easily identifiable directions, i.e., one that is parallel to the faceline at the center of the basin (CD in Fig. 2.1) and' the other is that perpendicular to the faceline but parallel to the diction of face advance (AB in Fig. 2.1). Nearly all the subsidence data obtained in the US have been derived from these two cross-sections, although some cross- sections parallel to CD but near the edges of the panel have also been included. In addition to moving horizontally toward the center of the basin, every point in the basin also subsides vertically. The magnitude of subsidence increases toward the center of the basin. Therefore surface subsidence is a three-dimensional problem and should be treated so in all cases. On all the major cross-sections, only principal subsidence and principal displacement occur. Since subsidence and displacement vary continuously in every major cross-section, three additional deformation components are de- rived, i.e., slope, curvature, and strain. On all other non-major cross-sections on the other hand the five components are accompanied by two additional components, i.e., twisting and shear strain. The seven components of the surface movement are defined as follows (Fig. 2.2): 1. Subsidence, S. On any cross-section, the vertical component of the surface movement vector is called surface subsidence. It generally points downward. But sometimes it points upward in areas ahead of the faceline or beyond the edges of the opening. In such cases it is a surface heave which is usually less than 6 in. 2. Displacement, U. On any cross-section, the horizontal component of the surface movement vector is called surface horizontal displacement. It generally points toward the center of the subsidence basin. But in steep terrain, it moves along the downdip direction 3. Slope, i. On any cross-section, the difference in surface subsidence between the two end points of a line section divided by the horizontal distance between the two points is called the surface slope of the section. 4. Curvature, K. On any cross-section, the difference in surface slope between two adjacent line sections divided by the average length of the two line sections is called the surface curvature of those two line sections. There are two types of curvature: con- vex or positive curvature and concave or negative curvature. 5. Horizontal strain, e. On any cross-section, the difference in horizontal displacement between any two points divided by the distance between the two points is called horizontal strain. If the distance between the two points is lengthening, it is tensile strain with positive sign. Conversely, if it is shortening, it is compressive strain with negative sign 6. Twisting, T. On the surface of the subsidence basin, the difference in slope between two parallel line sections divided by the distance between the two line sections is called twisting. 7. Shear strain, y. Shear strain is the changes in internal angles of a square on the surface of the subsidence basin or on any major cross-section. It is the summation of the differences in incremental (or decremental) lengths between the two opposite sides divided by the original distance between the two opposite sides. More precisely, the surface deformation indices (i.e., slope, strain, curvature, twisting and shear) are defined by derivatives of surface movement components. For simplicity, the x- and y-axes of the cartesian coordinate system are set to be parallel and perpendicular to the cross-section of interest, respectively. In such a coordinate system, slope and curvature along x direction are the first and the second derivatives of the vertical components (S) of surface movement with respect to x, respectively, or i, = ds/dx and kx = d2s/dx2. Horizontal strain along x direction is the first derivative of the component along x direction of the horizontal displacement,

Jan 1, 1992

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

By S. Horton, F. L. Stanley, P. J. Bethell

Introduction The McClure River Processing Plant is located at McClure in Dickenson County, VA. The plant was originally constructed in 1979 as a 600 st per hour facility. The plant is fed by coals from numerous mines in the area as well as a major longwall mine onsite. Coal is processed predominantly for the metallurgical coal market. The original plant was comprised of three separate processing circuits incorporating a heavy medium vessel for the coarse material (+10 mm). Intermediate sized material (10 x 0.6 mm) was beneficiated in heavy medium cyclones, with the remaining fines (0.6 mm x 0) passing to froth flotation for upgrading. The products from all three circuits were combined prior to thermal drying and then stored in silos or ground storage prior to shipment to the metallurgical coal markets. A middlings heavy medium cyclone circuit was also present to rewash the vessel and primary heavy medium cyclone rejects. The original circuit is as shown as Fig. 1 (Editor's note: All figures are reproduced at the end of this paper.) In reasonably close proximity to the McClure Plant is the much older, less efficient Moss 1 Plant, (Chance cone, table, water only cyclone, froth flotation, Fig. 2), which prior to the McClure Plant expansion was processing considerable tonnages of coal. At that time, the McClure River Plant was processing coal at the maximum possible feed rate, and the Moss 1 Plant was finding increasing difficulty in maintaining the stringent quality specifications on its major metallurgical coal accounts. Given the fact that many of the feeder coals to Moss 1 could be moved with minimal difficulty and cost to McClure, an exercise was conducted to evaluate the potential of expanding the McClure Plant. This would result in transferring coals from Moss 1, thereby increasing yield, reducing operating costs, and enabling customer specifications to be met more easily. After evaluating many possible ways of expanding the McClure Plant, we eventually settled on an approach based on the idea of Gallagher at German Creek in Australia (Gallagher, 1986). Our objective was to increase plant capacity by removing the coarser size fractions (0.60 x 0.15 mm) from froth flotation and at the same time eliminating the finer fractions (1.25 x 0.6 mm) from the heavy medium cyclone circuit, thereby allowing these circuits and therefore the plant as a whole to increase its capacity. These size fractions (0.60 x 0.15 mm from froth flotation and 1.25 x 0.6 mm from the heavy medium cyclone circuit) would be treated in a new spiral circuit. After desired evaluation of the various circuit capacities, including desliming screen, clean coal filter and thermal dryer capacities, etc., it was concluded that we should be able to maintain high levels of processing efficiency in the existing circuitry while at the same time expanding plant feed rate from a nominal 600 st per hour to 1000 st per hour, thereby achieving the objectives The circuit that was finally settled on is shown in Fig. 2. Several unique features were incorporated in this circuit in an attempt to maximize processing efficiency and reduce maintenance costs. First, the idea of a single-stage spiral circuit was abandoned in favor of a two-stage circuit incorporating rewashing the primary spiral middlings in a secondary spiral (Bethell, 88). The spirals selected were Mineral Deposits Model LD4 units. It was felt, after much previous testing, that spiral middlings rather than being a true middling gravity fraction were in fact a combination of true middlings with a preponderance of misplaced light coal and high-density rock. Consequently the insertion of this material directly into either clean coal or refuse streams (per normal practice) would only lead to loss of efficiency. The operating philosophy behind the design of the circuit was to produce the purest coal and rejects possible by adjusting the spiral splitters. Misplacement of refuse into the clean coal and vice versa would be minimized by taking a fairly large middlings cut. By reprocessing this material (10%-15% of primary spiral feed) in a secondary bank of spirals the misplaced material in the primary middlings would be reprocessed and its probability of reporting to the correct stream greatly enhanced. To minimize fines loading on the spirals, fine aperture sieve bends were to be included. A further novel feature of the circuitry is the desliming/ dewatering system. A philosophy of accepting inefficiency but forcing it to work in our favor was developed. The dewatering screens (Linatex Combis) were chosen not for their ultimate screening efficiency but for their ability to remove slimes (tolerating a fair quantity of oversize in the effluent) and their enhanced maintenance characteristics. Any oversize passing into the effluent, together with the fine coal centrifuge effluent, would pass to classifying cyclones, the objective of which was to reject to overflow (and hence froth flotation) as many fines as possible (-150 µm) without running the risk of losing coarse (1.25 x 0.15 mm) material to oversize. Hence, by a process of producing a "slime free" screen product coupled with a "coarse free" cyclone overflow, sizing efficiency could be maximized. Obviously, recirculating load had to be factored into the Combi Screen design capacity. The spiral circuit as shown in Fig. 2 was installed at McClure River in January 1988. The spiral circuit efficiency was very high immediately, and, after the normal debugging, plant capacity has been increased to the point at which we are currently processing of the order of 1100 st per hour. Test program The performance of the spiral circuit has been evaluated by two major samplings of the entire new circuit, coupled with routine bi-monthly tests on all other major components of the plant circuitry.

Jan 1, 1992