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By Francisco A. Fonseca

The Carajás Hills are in Brazil's Eastern Amazon, between the Araguaia and Xingu rivers, and parallels 5° and 7° Latitude South. A location map is given in Fig. 1. It is an area of low mountains (maximum altitude 900 meters), covered by the Amazon rain forest. Annual average temperature is 25°C, ranging from 31°C (maximum) to 19°C (minimum). Annual rainfall averages 2,000mm. Although it rains throughout the whole year, heavy rains occur from December through March. At the beginning of the 20's, Brazilian explorationists made brief geological descriptions of the area. Due to the thick soil, scarce fresh rock outcrops and the tropical forest covering a completely unpopulated area, detailed geological studies were difficult and reconnaissance expeditions were limited to the river margins.

Jan 1, 1982

By R. A. Garling

INTRODUCTION Since the infancy of in situ uranium mining, a growing number of hydrometallurgical processes have been incorporated into pilot and commercial scale flowsheets. Although initial design efforts were geared toward maximizing uranium recovery and minimizing plant and wellfield flow circuit maintenance, recent emphasis has shifted to improved means of water conservation and aquifer restoration. As mining units approached depletion, evaporation ponds reached minimum freeboard, and state and federal agencies demanded proof of groundwater restoration, processes including mixed bed and conventional ion exchange, reverse osmosis and electrodialysis were adopted by the industry. These units served the additional function of reducing process bleed flows during mining in states where the deep disposal well permitting ice remains unbroken. This report concerns the use of electrodialysis as an alternative to the more conventional processes used in in situ mining. In addition to a brief history and description of the process, a comparison to reverse osmosis and operational data derived from testing an Ionics, Inc. 1.31 x 10-3 m /s (30,000 gallon/day) unit at the Teton-Nedco Leuenberger Research and Development pilot will be presented. HISTORY Commercially practicable electrodialysis was contingent upon the development of synthetic ion exchange membranes in 1940's. In 1952, Ionics Inc. demonstrated that the process was amenable to the treatment of salt and brackish water and, in 1954, made their first commercial sale. The following decade saw several major electrodialysis unit sales which were generally targeted for use on private or municipal potable water treatment. Major increases in membrane desalting unit capacities, facilitated by technological advances in the reserve osmosis industry, were noted during the 1970's. The development of polarity reversing electrodialysis equipment which reduced feed pretreatment requirements, increased water recovery rates, and simplified unit operation, kept Ionics Inc. competetive in the water treatment industry. Engineering advances which incorporated automated equipment, non-corrosive construction materials, and improved ion exchange membranes allowed the electrodialysis process to compete in industrial waste treatment among other commercial markets. PROCESS AND APPARATUS DESCRIPTION The electrodialysis process utilizes direct electrical current passed across a stack of alternating cation and anion selective membranes in order to achieve an electrochemical separation of ionized materials in an aqueous solution. The membrane stack has the appearance of a plate and frame filter press and auxilliary equipment includes solution pumps, electrically actuated valves, filters, piping and a direct current power source. The ion separation membranes are thin sheets of synthetic cation or anion selective resins. Attaching sulfonate or quaternary ammonium groups to the cross linked copolymer structure determines the ion selectivity of the membrane. The membranes are separated from each other in the stack by non-conductive spacers that house flow channels which route the flow tortuously and parallel to the membranes. Direct electrical current passing perpendicularly to the membranes and solution passages attracts cations toward the cathode and anions toward the anode (Figure 1). As the ions from the feed stream pass through the ion selective membranes, they become concentrated in the adjacent brine channel and are retained there by the combined attractive force of the electrode and the repelling force of the next membrane toward the electrode. Limiting factors on the degree of demineralization possible include chemical solubilities in the brine flow and the current density that will produce an unacceptable degree of polarization (Figure 1). Feed or brine solution treatment with complexing agents or acids has been successfully applied to prevent membrane scaling. Polarization can occur when sufficient current density is applied to dissociate water in the ion depleted region of the diluting compartments near the membrane surfaces. Significant polarization is evidenced by large electrical resistances across cell pairs and notable pH differences between diluting and concentrating streams. Limiting current densities have been increased in U.S. manufactured equipment by utilizing tortuous flow paths of relatively high linear velocities thereby promoting continous solution mixing. Energy consumption is due to separating electrolytes and solutions, oxidation and reduction reactions occurring in electrode compartments, overcoming electrical resistance, conversion from AC to DC power, solution pumping and auxiliary equipment actuation. A major improvement to the basic electrodialysis process was applied in 1970 which resulted in frequent, automatic cleaning and descaling of membrane surfaces. The process, polarity reversal, incorporates alternating the cathode and anode on a periodic basis while exchanging product and brine flow channels via electrically actuated values. The reversal reduces the potential of stack plugging with CaCO3 (calcite), CaSO4 (gypsum), and colloidal materials and, in most waters, eliminates feed pre-treatment requirements. For approximately two minutes during and following the reversal, off spec. water is flushed to waste or reintroduced to the feed supply. The usual feed treatment on polarity reversing electro-

Jan 1, 1982

By Daniel R. Walton

This paper covers several points of interest to mine engineers and planners involved in evaluating or conducting geostatistical coal reserve estimates. In particular, the ability to optimize borehole drilling locations and minimize drilling costs is addressed. Also discussed is the concept of confidence intervals for estimated coal characteristics obtained through geostatistical estimates and their importance in decision-making for mine planners. Finally, a discussion is presented on how geostatistics enables identification of invalid data points and actually forces strict attention to data editing and proper geologic interpretation.

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

This paper discusses some topical issues on the mining administration and legislation fields including:  1. Licensing û Administration and associated specialist functions including:Processing and volumes of mining privilege applicationsFuture organisation and possible decentralisationSpecialist staffFacilitiesGuide to the Mining Act 1971 2. Mining Amendment Act 1981 Transitional procedures 3. Planning Tribunal First hearing of objectionsPoints of principle resulted there from 4. Quarries and Tunnels Bill Currently before Parliament û implications for metallurgical mining

Jan 1, 1982

With the proclamation of the Environmental Protection Act, 1971-80, the Environmental Protection Authority and Department of Conser- vation and Environment were established to enhance the quality of the environment. Proposed gold developments of all scales can be referred to the Authority in the form of a Notice of Intent. Environmentally significant projects might then require preparation of an Environmental Review and Management Programme. Current environmental issues include the'effects of clearing by prospectors,, mine dewatering and tailings disposal. INTRODUCTION In the past, the environmental aspects of gold industry related activities have received little attention. However, with the proclamation of the Environmental Protection Act, 1971-80, mechanisms were instituted which allowed both for referral of projects to the Environmental Protection Authority (EPA) and establishment of the Department of Conservation and Environment. As well as providing assistance to the EPA on referred projects, the Department is concerned about a number of environmental issues associated with the gold mining industry.

Jan 1, 1982

By William H. Mount

INTRODUCTION Effective mine ventilation is required to maintain a healthy underground environment for humans. Without effective ventilation, the environment can become unhealthy or hazardous as a result of the depletion of oxygen, contamination with toxic gases, or the buildup of an excessive amount of particulate matter (dust). Each contaminant has an upper limit of concentration that should not be exceeded within an 8-hr period. This is known as the threshold limit value (TLV). The TLV represents an acceptable level of exposure that should produce no ill effects. Unfortunately, more than one contaminant may be present at any one time and the effects of the individual contaminants may be additive, i.e., the effects of each contaminant must be considered simultaneously to determine the potential danger to miners. Possibly the greatest threat to mine-air quality is the uncontrolled underground use of the diesel engine. The diesel engine was invented in 1892 by Rudolf Diesel. His intention was to develop a power source that could burn coal dust as a fuel, but he was unsuccessful in that attempt and had to resort to liquid petroleum fuels (Johnson, 1975). In 1898, the diesel engine was intro¬duced into the United States by Adolphus Busch, who anticipated using it as a prime mover in factories and generating plants. At that time, the diesel engine was a very large and very heavy engine, designed for fixed installations. By 1919, lighter engines were being developed, and by 1931, Caterpillar was marketing a diesel-powered, track-type tractor (Henderson, 1975). Since 1931, the diesel engine has evolved into an extremely popular prime mover in medium and heavy-duty applications. Efficiencies now are on the order of 40% and improvements such as turbocharging and aftercooling have produced engines capable of generating power at a ratio of 0.3 kW/kg (1.0 hp per 5.0 lb) of engine weight. Although the diesel engine is relatively efficient as a mobile power plant, it is far from efficient in terms of the energy produced from the energy potential of the fuel. About 60% of the heat value of the fuel leaves the engine as wasted heat, with about 50% of that heat being emitted through the exhaust pipe and the other 50% being emitted through the radiator or cooling fins. Perfect combustion in an engine would produce only water vapor (H_0), carbon dioxide (CO_), and nitro¬gen (N2) as the byproducts. Since the diesel is not a perfect engine, each pound of fuel burned generates 5.6 m° (200 cu ft) of exhaust gas, containing about 0.009 m3 (0.33 cu ft) of carbon monoxide (CO), 0.009 m' (0.33 cu ft) of nitrogen oxides (NO, NO2, and NOD), and 0.57 m3 (20 cu ft) of carbon dioxide. The balance of the exhaust emission consists of free nitrogen and water vapor (Hurn, 1975). Contrary to popular belief, the diesel is not inherently dirty. Under normal operating conditions, a well-maintained engine neither smokes nor smells. However, the same well-maintained engine can and does produce toxic emissions. A diesel engine is not inherently safe and constitutes a distinct hazard to personnel. This chapter is devoted to a description of the various toxic substances that may be generated by a diesel engine. Although some of these substances may also be produced by blasting or natural causes, the focus of this chapter is on the relationship between the internal- combustion compression-ignition engine (the diesel) and the quality of the mine air. CARBON MONOXIDE Combustion Process During the combustion process (burning) of organic fuels, each atom of carbon combines with two atoms of oxygen, provided that a surplus of oxygen atoms is available. Thus, the carbon is oxidized to carbon dioxide. Most open flames, such as trash fires, camp fires, gas ranges, etc., produce carbon dioxide. However, with insufficient oxygen, incomplete combustion results as the carbon atoms each combine with one atom of oxygen to produce toxic carbon monoxide. Burning charcoal briquettes produce carbon monoxide because the combustion takes place inside the briquettes where sufficient oxygen is not available to the combustion process. Internal-combustion engines, whether burning gasoline or diesel fuel, also produce carbon monoxide. The only oxygen available to the combustion process is that trapped within the cylinder. If the amount of fuel delivered to the cylinder is excessive, there is insufficient oxygen for complete combustion and carbon monoxide production results. In a normally aspirated (nonturbocharged) diesel engine, the amount of air "sucked" into a cylinder is the same on every intake stroke, resulting in complete combustion only at low levels of engine loading, when small amounts of fuel are injected. Higher levels of engine loading cause larger amounts of fuel to be injected into the same volume of air in the cylinder. Unless the volume of air is increased, the combustion process becomes progressively less complete as the amount of fuel increases. Turbocharged engines are able to compensate some¬what for increased loading and increased fuel consumption. The turbocharger acts as a compressor for the intake air, forcing a larger volume of air into the cylinders as the engine speed increases. Hence the turbocharged engine burns cleaner than a naturally aspirated engine and produces slightly less carbon monoxide (Marshall and Fleming, 1971). Much of the underground equipment used today is powered by turbocharged indirect-injection diesel engines. Although these engines emit fewer toxic contaminants than naturally aspirated engines, they do not eliminate the problem. Since the turbocharger is driven by the exhaust gases, rapid accelerations can cause temporary overfueling of the engine until the turbocharger attains a speed sufficient to restore the correct air-to-fuel ratio. During this "turbocharger lag," the combustion cylinders contain insufficient oxygen, causing severe smoking and an increased output of carbon monoxide.

Jan 1, 1982

US 4,190,434 - In the thermal production of magnesium metal, a mixture of calcined dolomite, an iron- silicon-aluminum alloy as a reductant, and residual slag from the production of ferrochromium is smelted to evolve magnesium vapor, which is then condensed. The use of bauxite is lessened or eliminated because of the relatively high alumina content of the residual slag. R. Bonfils, A. Mena, C. Payn, and L. Septier, assigned to Ste. Francaise d’Electrometallurgie (SOFREM). Feb. 26, 1980. 6 pp. (75-10R). Filed: 6-14-78 (915,387). Priority: France, 6-24-77 (77/20,227).

Jan 1, 1982

By Sandip K. Mukherjee, Anthony W. Laurito, Madan M. Singh, Jon C. Volkwein

In the past, little attention has been paid to the costs of ventilation, since it seldom represented a significant portion of the total mining cost. However, in recent years adverse mining conditions and increases in power charges have caused ventilation costs to rise. The federal regulations have set minimum standards for volumetric airflow rates in underground coal mines. As a mine develops, the air volume and pressure requirements increase. There are several methods by which the quantity of airflow to the mine and working faces can be increased. Some of the methods include (1) increasing main fan output or replacing the main fan by a new higher capacity fan, (2) reducing leakage across stoppings and overcasts, (3) adding a new ventilation shaft or borehole, (4) reducing resistance to airflow by reducing the resistance factor of the individual airways, increasing the number of airways (especially the parallel airways), or re- arranging the air courses by using the available entries, and (5) using auxiliary ventilation systems. This paper addresses the cost-benefit analysis of several of these alternatives as applied to a large operating coal mine located in southern Illinois. Pressure and quantity surveys were conducted under-ground, and measurements were taken at the main exhaust fans on the surface. The existing ventilation network was modeled by using the Penn State ventilation simulator. The input data included the configuration of branches, resistance of each branch (including a component for shock losses), and changes in pressure and volume. The results of the simulation were compared with the survey data to check the merits and accuracy of the model network. The net- work was designed to take into account shafts, major entries, overcasts, regulators, gob areas, working faces, and major leakage paths. The model network served as the basis for the cost analyses of networks designed for increased airflow. The applicable alternatives for increasing airflow were simulated. Performance data obtained from these simulation runs were utilized to conduct the cost-effectiveness analysis. The costs for each f the alternatives included capital, operating and maintenance costs. These costs were expressed as annual costs in terms of dollars per cfm of circulating air, taking into account the economic life associated with each of the alternatives. The benefits were expressed as increased airflow quantity in the critical sections. The result of these analyses show that for the short-range plans evaluated, the most cost-effective way of increasing airflow at this mine would be to reduce leakage through stoppings and overcasts.

Jan 1, 1982

By S. S. Peng

This paper presents a convenient way for the preliminary design of the shield supports. The locus equations which are applicable for all kids of shield supports are derived. These equations facilitate the statics analysis of the shields. An indirect method in determining the coefficient of friction, the external toads and their locations are also developed.

Jan 1, 1982

By Richard K. Thorpe, Gordon D. Anderson, Merle E. Hanson, Ronald J. Shaffer

The propagation of pressurized fractures across a frictional interface is discussed, with emphasis on the case where an offset, or step, in the crack is produced. Theoretically, the steps can occur at regions of reduced shear strength along the interface. As a fracture is propagated toward a weak zone, extensional strain is concentrated at the edges of the zone where higher shear stresses can be sustained. Thus, when the fracture intersects the weak zone, it can be reinitiated at some small distance away on the opposite wall of the interface. The phenomenon has been studied through numerical modeling and laboratory experiments of the hydraulic fracturing process. Field observation of natural step cracks indicates that the mechanism is applicable to the genesis of jointing patterns in rock. Such features should not be confused, therefore, with steps caused by shear displacement of an interface.

Jan 1, 1982

By John A. Rajala

The water chemistry of the froth flotation pulp is one of many parameters which can affect the flotation process. The fluid may contain certain inorganic salts which, along with increased ionic strength, can produce undesirable flotation depression or activation. A source of a salt could be the use of calcium chloride to prevent freezing in ore cars, chutes, and bins (Anon., 1980). Alternately, the salt or salts could be initially contained in the water as in the case of sea water and some brackish waters. These salts may become more concentrated because of the need for recycling of plant water due to environmental regulations and/or economic considerations (Davis, F.T., et. al., 1976). The presence of inorganic salts in flotation pulps and their effects upon the recoveries and surface chemistries of different oxide minerals have been the object of various studies. Some researchers (Fuerstenau, D. W. and Raghavan, 1976; Fuerstenau, D. W. and Modi, H. J. 1960; Kulkarni, R. D. and Somasundaran, P., 1977) have reported that addition of a salt can increase or decrease flotation recovery depending upon the particular system.

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION This method is a variation of cut-and-fill mining where either the wall rock or the vein is too unconsolidated to permit safe mining by conventional cut-and-fill methods. Basic development consists of driving a 2.4 x 2.7-m (8 x 9-ft) haulage heading laterally in the footwall side of the vein. Crosscuts are driven over to the vein at 91-m (300-ft) intervals. Then the vein, at each inter¬section, is mined out to a distance of about 9 m (30 ft ) above the track level. After the installation of an ore chute, scrams or intermediate drifts are driven in both directions along the vein. Mining then starts imme¬diately below the level bounding the upper limit of the stope. A cut varying between 3 and 4.2 m (10 and 14 ft) in depth and mined for a distance of 45.7 m (150 ft )along the strike is taken from the orepass in each direction (Fig. 1). In this mined-out void, the floor of the cut is equipped with 203 x 203-mm (8 x 8-in.) caps hitched into the wall with steel saddle hitches installed on 1.5-m (5-ft) centers. Double wood stringers, 152.4 x 152.4 mm (6 x 6 in.), are then installed on 1.2-m (4-ft) centers longitudinally down the length of the stope. Reinforc¬ing wire mesh is also placed, overlying each mat of timber. Directly above the stringers, stulls are installed from hanging wall to footwall to hold the timber in place during the sand-filling operation. Then a cement¬sand mixture in a ratio of about 1:10 weight is poured to a depth of about 2.4 m (8 ft). The entire cycle is then repeated down to the scram drift. Another variation of timbering is illustrated in Fig. 2. There is a vein width of approximately 5.5 m (18 ft) with the same unconsolidated ground conditions. The primary difference between handling this situation and the previously described undercut-and-fill method is that standard timbered sets are installed in each excavated stoping void rather than stulls. Once again the basic method is to work down the ore block from an upper timbered drift to a lower timbered drift. Cuts are taken from a timbered raise with a chute on the stope side. Each piece of ground excavated is essentially two sets high, 5.5 m (18 ft), by two sets wide, 5.5 m (18 ft), by the length of the cut along the strike of the vein. The timbered sets are 2.4 x 2.4 x 2.4 m (8 x 8 x 8 ft). The net result is, therefore, two sets of timber on a top floor separated by doubled-up 76-mm (3-in.) lagging from

Jan 1, 1982

By Joseph W. Murray

GENERAL DESCRIPTION The undercut-and-fill method of mining was devel¬oped by Inco Ltd. in the Sudbury district of Ontario, Canada, to deal with abnormal ground conditions en¬countered in pillar recovery. Although it is now some¬times utilized as a primary stoping method, the main application continues to be in pillar recovery. In any mining method in which fill is employed, pillar systems tend to fail gradually as stoping pro¬gresses. As a result, most pillars are completely destressed and broken by the time pillar recovery com¬mences. Control of the back in such ground becomes a major problem and is contingent on local conditions as they exist in the workplace. In most mining methods, operations such as drilling, blasting, slushing, and tim¬bering can be readily assessed and the mining cycle pre¬dicted. However, in the mining of pillars with overhand methods, the time and effort required to maintain a safe back varies to a quite unpredictable extent. For years the square-set method was the most effective means available for pillar recovery. However, in areas where bad ground was encountered, square setting was a costly and sometimes hazardous method. As the tonnage mined from pillars increased, it was apparent that a safer and less expensive approach was required. Past experience indicated that the best way to deal with broken ground was to mine from the top down. Top slicing systems were examined but were rejected due to the major sub¬sidence which resulted. However, ideas generated from studying top slicing finally developed into the present method of undercut-and-fill. This method, which has almost completely replaced square-set mining, exhibits reduced costs, has improved safety and efficiency, elimi¬nates the factor of bad ground overhead, and allows a closer prediction of the actual mining cycle. The result is more accurate scheduling of pillar production. This factor is of increasing importance as the tonnage from pillar recovery increases in relation to total mine pro¬duction. Undercut-and-fill is a method of extracting a block of ore by mining successive cuts, working from the top down. After a cut of ore is completely mined out, lami¬nated timber stringers are constructed along the sides for the full length of the cut. Round logs covered by a wire screen are laid across the stringers to form a mat. The cut is then tightly filled with hydraulically placed cemented sand fill. Mining is then resumed on the cut immediately below the mat. As the cut advances, the timber stringers are supported by round timber posts seated on the solid bottom of the cut. Drilling, blasting, muck removal, and timbering operations are repeated until all the ore in the cut has been removed. At this time, another log mat is laid and the cut is filled. This sequence is continued until the block has been mined from level to level. The basic principles of undercut-and-fill can be adapted to a variety of mining situations. It can be applied transversely or longitudinally in ore bodies of varying width, that dip from vertical to flat lying. The method can be applied using conventional equipment or it can be mechanized through the introduction of small diesel or air-powered units for muck removal. Undercut¬and-fill is a selective method, ore recovery is high, and dilution can be controlled to an acceptable level. How¬ever, it is a high-cost, labor-intensive method and as such its use is restricted to areas where the less costly meth¬ods of ground support are not successful or where bulk mining methods are not adaptable. One basic require¬ment for undercut-and-fill is an adequate source of hydraulically placed cemented sand fill and a dependable sand plant and distribution system. The success of the undercut method is to a large degree dependent on the competence of the fill placed in each of the successive cuts. In theory, the weight of the fill on each cut is supported in part by the cemented sand itself as it binds to the rough sides of the mined¬out areas next to the pillar walls, and in part by the tim¬ber support. As mining progresses beneath the mat, a downward pressure is exerted within the cemented fill and a voussoir arch develops above the mat, across the full width of the cut. This supports the major portion of the load. The log mat and posts have a double func¬tion in that they carry the remaining weight not sup¬ported by the voussoir arch and in addition prevent any pronounced settling of the fill mass. Excessive settling breaks the bond between the cemented fill and the pillar walls, eliminating the arch effect and transferring the entire load to the mat and posts. The theory of a voussoir arch forming within the cemented sand fill mass is substantiated in practice by the fact that pillars have been mined from level to level with no more significant weight on the bottom cuts than was experienced on the top cuts. In effect, as mining progresses downward, there is no additional transfer of weight from previous cuts to the cut being mined. APPLICATION OF UNDERCUT-AND-FILL Because of their size, the Sudbury ore bodies are, for the most part, mined transversely and, to a considerable extent, by filling methods. Transverse stoping blocks are designed with 9-m (30-ft) crown pillars established be¬low each level and with stopes separated by vertical rib pillars. In past years, floor pillars were left above the level as well, but present practice favors the silling of primary stopes at the base of the rail. Rib pillars are designed to meet the ground control requirements of the primary stoping methods. Local conditions and past experience generally dictate their size. Historically, since the square-set method was to be used for recovery, rib pillars were designed to accommo¬date standard 1.7-m (5'h-ft) sets. A typical rib pillar is four sets wide [6.7 m (22 ft)] when located between two square-set or cut-and-fill stopes having a width of 8.4 m (27.5 ft). When the undercut-and-fill method was de¬veloped, a mat width of two sets was adopted. This di¬vided each cut of the common four-set pillar into two

Jan 1, 1982

By V. J. McSweeney

This paper is directed at interested parties outside the Beach Sand Industry : the scope is quite broad. The geomorphology described is a successful working hypothesis which guided the author during prospecting and pegging campaigns : the author is not an expert in this field. The test data used in the text, for illustrative purposes, is original and may constitute a contribution to the state of the art. The operation- al and design philosophy embodied in the text is personal and describes the author's methods : all practising exponents would differ in some degree amongst themselves on these subjects. In this paper, no criticism of any existing operation or brand of equipment is intended and names are used to maintain a real contact with the subject matter. The author believes that the operational and design techniques evolved in the Beach Sand Industry under continual duress, have a wider application in the mining industry. This glimpse of the Industry is intended to enhance the respect for the present and past members. Space does not permit individual acknowledgements.

Jan 1, 1982

Continuous remote sensing of air speed In underground mine airways. Approach An ultrasonic technique measures a form of air turbulence which is proportional to air velocity. How It Works The stainless-steel enclosed transducer consists of a sensing head with circuitry, and a telemetry card. The sensor uses a patented ultrasonic technique to measure a form of turbulence created In the air. This turbulence, known as vortex shedding, is related to the air speed through the sensor. Air flowing past a cylindrical post creates vortices which flow down the air stream on both sides of the post. Pulses from an ultrasonic transmitter pass through the vortices to a receiver; the effect of the vortices on the pulses is a measure of air velocity. The transducer contains a telemetry card, known as an accessor, which transmits reports on air velocity through a pair of wires to a monitoring system. The accessor contains an analog-to-digital converter and digital telemetry circuitry required for computer system operation. The transducer does not contain any read-outs or alarms and air velocities are displayed by means of a centrally located

Jan 1, 1982

By B. J. K. Sullivan, Brian Hornsby

INTRODUCTION Mount Isa Mines Ltd. operates a copper and silver- lead-zinc property at Mount Isa in northwest Queens- land, Australia. Mount Isa is located in the tropics at latitude 20°43'S and longitude 139°30’E (Fig. 1). It is on the upper watershed of the Leichhardt River and at an elevation of 365 m above sea level. The average rain- fall is 360 mm/y, the greater part of which falls in January to March (the hot summer months). The mean summer temperature for these months is 30°C (dry bulb), with a maximum temperature of 43°C; the mean wet bulb temperature is 25°C with a maximum of 28°C. Mount Isa is located geologically towards the western edge of the Precambrian Shield of northwestern Queensland. This complex forms a north-tilting dissected peneplain which extends in the north-south direction for 640 km, with widths varying from 32 km to 190 km. The combination of differential erosion, regional strike, and steep dips has produced a corrugated topography of low north-south orientated hills and valleys. The field was first discovered in 1923 as a silver- lead-zinc field. Mount Isa Mines Ltd. was formed the following year and subsequently became the sole operating company on the field. Lead and copper ore have been mined and treated separately and simultaneously at Mount Isa since 1953. Intensive exploration commenced in 1954 and an active exploration program provided the justification for a series of production expansion programs in both silver- lead-zinc and copper (see Table 1 ). The most recent and largest production expansion program at Mount Isa commenced when a decision was reached in December 1969 calling for an increase in copper production from 100 000 to 150 000 t/y of blister copper. A major factor which influenced this decision was the substantial additions to the ore reserves in the so-called 1100 ore body which is a massive copper sulfide ore body with a known strike length of approximately 2500 m, a height which is variable up to 300 m, and a width of generally 300 m up to a maximum of 370 m (see Figs. 2 and 3). Host to the copper mineralization is a complex and variable rock generally called "silica dolomite." The ore body strikes 6.02 to 6.1 rad (345° to 350°), dips west between 0.78 and 1.13 rad (45° and 65°), and is truncated at depth by fault implaced basement rocks. It first occurs some 600 m below the surface and bottoms at about 1000 m below the surface. The unconfined compressive strength of the silica dolomite ore varies from 150 to 250 MPa, whereas the basement rocks, carbonaceous mylonite, slaty shale, buck quartz, and greenstone (a chloritic schist) are considerably weaker, varying in strength from 20 to 80 MPa. The distribution of mineralization within the ore body is variable, but can be considered as generally high grade at or just above the basement contact and diminishing towards the upper reaches of the ore body. The current

Jan 1, 1982

By Alan Rhodes

This study is initiated to investigate the efficiency of separation of mineral slimes using electroflotation as compared to that of column and conventional flotation techniques. Important parameters such as collector concentration, pH, current density, aeration rate, etc., have been studied. The results indicate that electroflotation can be used more efficiently in processing rutile slimes (10 µ) under certain operating conditions.

Jan 1, 1982

By Michael J. Pratko

The bulk handling industries have long been plagued with problems in their attempt to obtain one of their major goals - the maximum efficiency for which a bulk handling flow system was designed. Mentor Dynamics Ltd. is an international firm with offices in Canada, Japan, and the United States. Mentor Dynamics Ltd. is in the business of solving the "why not" of a system and engineers the solution for the problem. Studies of the varied difficulties encountered with bulk flow systems indicate that 90% of all flow problems fall into one or more of the following categories: abrasion, impact, sticking, and/or inadequate design. These problems can occur with a myriad of bulk solids: coal, gypsum, aragonite, phosphate, fly ash, ore. Solutions to the problems are as varied as the problems. The proper lining material, ceramic, urethane, rubber, and/ or ultra high molecular weight polyethylene, must be coupled with the correct fastening system, annular ring and 'F' cap, weld washer, button head fastener, weld studs and/or nuts and bolts.

Jan 1, 1982

By J. J. Pavlica

Grinding in steel mills adversely affects the flotation removal of pyrrhotite from magnetite ores. To elucidate the effect of grinding media on the floatability of pyrrhotite, rest potential and short-circuit current measurements were made on electrode combinations of pyrrhotite-mild steel and pyrrhotite-stainless steel. Galvanic contact of pyrrhotite with mild steel increased the rate of surface reactions on pyrrhotite in air-exposed and oxygenated solutions. It was postulated that the formation of ' an iron hydroxide coating on the pyrrhotite adversely affected its floatability. Its formation was minimal when pyrrhotite was contacted with stainless steel. Fine magnetite particles caused slime coating on pyrrhotite surfaces by magnetic interaction and impaired the flotation of pyrrhotite.

Jan 1, 1982