Search Documents

By G. Bhaskar Raju, S. Prabhakar

The amenability of column flotation for the beneficiation of fluorspar was investigated, and the overall performance of the column, flotation was compared to the performance of the existing conventional flotation cells at the plant. It was estab¬lished that improved grades and recoveries could be obtained by adopting column technology. Acidspar concentrates as¬saying 97% CaF2, suitable for HF manufacturing, can be produced by adopting two-stage column cleaning in place of multistage cleaning by conventional cells. Similarly, metspar concentrates assaying 90% CaF2, suitable for metallurgical applications, can be achieved in a single-stage cleaning by column flotation. This is an improvement over what is now being achieved using the existing conventional rougher and three-stage cleaning cells in the plant. The number of stages, i.e.,four in the acidspar circuit and four in the metspar circuit, including rougher stages, can be eliminated by adopting a three-column configuration. Thus, the cost of production, the complexity of the circuit and the power consumption can be effectively reduced with other attendant benefits.

Jan 1, 2000

By Robert F. Randolph

New technologies confront mining organizations with unprecedented human factors challenges. Likely problem areas range from the design of the human- machine interface to the design of work groups and mine organizational structures. However, these human design considerations are commonly ignored in most industries, resulting in a failure rate of 75% for advanced technology programs. The U.S. Bureau of Mines is addressing these concerns through a program of research on the human side of high technology mining systems. Interviews with mining industry participants and researchers have identified concern in several areas, including training and equipment design. Although new mining technologies promise to improve overall safety by relocating miners away from hazardous areas, data from MHSA accident reports and other sources suggest that certain types of new hazards may be created. Many of these new hazards can be reduced though a systems approach to ergonomic design of equipment, tasks, and work processes. This evidence, combined with reviews of high technology human-machine systems in other industries. can help in the design of safe and efficient future mining technologies.

Jan 1, 1992

By Firoze E. Katrak, John R. H. Black, Paul R. Ammann, Michael J. Loreth, Servi. Italo S.

Germanium is produced primarily as a by product from zinc concentrates. The important sources of gallium are alumina ores (bauxites) and zinc concentrates and residues from zinc processing. In addition, gallium as well as germanium may soon be recovered from a depleted copper mine near St. George, Utah. The major uses for germanium are in semiconductors, lenses in infrared optical systems, catalysts, and gamma radiation detectors. The major markets for gallium are in light-emitting diodes, solar cells, laser diodes, integrated circuits, microwave devices, and magnetic bubble memories. CRA projects a healthy growth in the demand for both metals. However, the role of scrap metal will continue to be important over the next decade, thus decreasing the overall demand for primary gallium and primary germanium. This paper summarizes the key elements and the economics of the markets for germanium and gallium. Much of the information, such as plant-by-plant costs of production and estimates of consumption by end use for various countries is from proprietary databases of Charles River Associates.

Jan 1, 1986

By P. C. Cochrane, I. S. Clelland, S. H. E. Phillips

Geological and hydrological conditions in the strata overlying the Salado Formation vary considerably in the Carlsbad, NM, area. The first four shafts sunk at International Minerals and Chemical Corp. (IMC) property have been located in competent and essentially dry conditions, while other shafts on adjacent properties have had such adverse conditions that ground freezing has been chosen as the water and ground control technique for shaft sinking. At the location of the IMC No. 5 shaft there has been extensive erosion and salt dissolution in the Rustler Formation, leading to collapse of some zones of the strata. Both ground freezing and pregrouting were considered as alternative methods of stabilizing the extremely poor ground conditions and controlling the potentially very heavy water inflows. It was decided that pregrouting the shaft was feasible and would offer substantial cost and time saving over a ground freezing technique. This paper describes the pregrouting techniques used and illustrates the requirement for flexibility in a grouting program so that it can be modified to reflect any new or additional information obtained during the progress of the grouting.

Jan 1, 1987

By L. Shevenell

Nevada now has, or will have, 35 pit lakes resulting from open-pit mining below the water table. State regulators are requiring that the ultimate pit lake water quality be predicted as part of the permit application process. Water balances/pit lake filling rates must be esti¬mated in support of the geochemical modeling. In Nevada, analytical and numerical methods have been employed for the study of pit lakes. In the current work, analytical solutions are investigated for one pit lake, the Getchell south pit lake, which previously filled be¬tween 1968 and 1982 during a hiatus in mining. Water levels vs. time are available for this lake and these data have been fitted with a hyperbolic curve. This water-level-vs.-time curve was used in simple water-balance computations to find the net inflow rates by forcing computed water levels and volumes to match those of the known water levels and volumes through time. The model described here provides realistic volume fluxes into and out of the lake on a monthly basis as the pit lake filled.

Jan 1, 2000

By Randal J. Barnes

In, any open pit mine the ability to make accurate production estimations for the shovel-truck system is an important asset. This is true in an operating mine wishing to estimate production under a set of system modifications (i.e. more trucks, fewer loaders, change in roads, etc.), as well as in the pre- mining planning stages of a property. Many different methods have been proposed for solving this problem; some of the methods rely heavily on "rules of thumb" while others are highly mathematical and require access to a high speed digital computer. Analysis of shovel-truck systems with the use of simulation models is a widely accepted technique within the industry. Through the use of such models it is possible to account for much of the inherent variability of the system and thus produce meaningful results. However, simulation studies are expensive. Formulation and setup of the actual simulation model requires a significant commitment of time and energy. In addition, the cost of the necessary computer runs is often of considerable magnitude.

Jan 1, 1979

By Louise H. Michaud

An interactive mine planning system that employs recent advances in 3D computer graphics capabilities and 3D geologic modelling to aid with the tasks involved in producing dragline mine plans was developed. As well as offering large potential saving in the time required to carry out the planning tasks, the system offers immediate visual feedback to the planning engineer. This can greatly aid in the conceptual understanding of complex geology and mine plans. The package incorporates the generation of a three dimensional Gridded Seam Model (GSM) as the basis for the development of the geologic database. This modelling method is capable of accurately representing rugged topography and multiple coal seams, including folding and faulting. The GSM is then integrated with an interactive graphics mine planning system in order to develop one model and one database for the generation of the mine design, the depletion of the coal resources and the movement of spoil material. The gridded structure of the model also provides the foundation on which all mine and spoil pile volumes are calculated.

Jan 1, 1992

By Mark A. Alspaugh, E. A. “Bud” Viren

The driving force behind most advances in technology is need. In the early 1980s, underground coal miners in the USA wanted to build longer conveyors to handle larger capacities but continue to use fabric reinforced belting with mechanical fasteners. Conveyors with intermediate drives were certainly the answer but achieving reliable designs required a significant learning curve. Basic design methodologies had to evolve including advanced simulation of starting and stopping along with complex control techniques and logic. Some components also evolved as specific performance requirements changed Today, in addition to underground mining, the technology has been used extensively in the tunneling industry where the need is for control of belt tensions to negotiate tight horizontal curves. Even though intermediate drives are not used extensively on the surface yet, the lessons learned will find their way to long overland surface conveyance where smaller components are more available and reliable, and often with less capital cost. This paper will document the evolution of intermediate drive technology over the last 30 years and will show examples of its current use.

Jan 1, 2004

By Joseph E. Worthington

The Haile Gold Mine is one of the earliest gold discoveries in the eastern United States as well as one of the largest gold producers from that region. Gold was first discovered in the gravels of what is now known as Haile Gold Mine Branch in 1828 on lands then owned by Colonel Benjamin Hale (spelling not the same). Small placer workings were immediately started and it was not long before the surrounding residual soils and saprolites were also found to contain gold. Following the presence of gold in the residuum eventually lead to the discovery of the underlying gold-bearing lodes that actually outcropped in saprolite form where not masked by heavy vegetation or a thin veneer of Coastal Plain sand. The earliest attempts at production, immediately after discovery, utilized slave labor, washing gravel and residuum from leased tracts that were 15m (50 ft) square. Several at- tempts at small milling operations were made during the en- suing decades, but there was really no substantial or successful production until the 1880s.

Jan 1, 1990

By Jinrong (Patrick) Zhang

Improvement of conventional flotation efficiency, separation of carbonates from apatite, recovery of phosphate values from phosphatic clay and other ways of utilizing the clay, disposal of clay slurries and subsequent reclamation of the settling areas, and conservation of resources and energy remain major challenges in the beneficiation area for the phosphate industry. These problems have been actively investigated by the industry, government agencies, and research organizations for decades with modest success. However, some recent research activities have generated promising results and possible new technologies. The dominant position of the Crago process in phosphate beneficiation is being challenged by some of the single stage processing schemes. Heavy media separation enables one producer to process a portion of high carbonate ores. Separation of dolomite from apatite by reverse flotation of carbonates with phosphate depressants showed potential for industrial application. Consolidation of phosphatic clays in mine cuts using a modified "sandwich" process is being tested. Beneficiation in the acidulation plant or grinding at the mines may be considered in the future.

Jan 1, 1993

In this presentation, CiDRA?s patented technology platform and its applications will be described. CiDRA's non-invasive, passive sonar array-based flow meter technology provides the volumetric flow rate of single or multiphase fluids by measuring the speed at which naturally occurring structures such as turbulent eddies or density variations convect with the flow past an axial array of sensors. These sensors are incorporated in a band that is wrapped around the outside of the pipe, resulting in no process downtimes for installation and unprecedented reliability. This technology has resulted in a unique ability to measure the flow rate of most fluids ? clean liquids, high solids content slurries, pastes, and liquids and slurries with entrained air. Unique and difficult minerals processing flow measurement problems and their solutions will be described. Examples of these problems include flow measurements in froth lines and flotation feed lines with entrained air, slurry lines with magnetite and other magnetic ore, slurry lines with abrasive or corrosive materials, high pressure lines, and lines exhibiting scale buildup. Recent technology developments will be covered.

Jan 1, 2008

By J. Villeneuve, M. A. Reute, E. Verhoe

The objective of EU waste policy is to reintroduce used materials into the economic cycle, especially by recycling, or to return them to the environment in a useful state or at least a harmless state. To support decision-making in complex systems, the standardised ISO 14040 life cycle assessment (LCA) method can be used to quantify and to interpret the environmental repercussions of (changes in) a product system from cradle to crave related to a functional service unit. The use of the standardised LCA method at European level can be characterised as a damage-oriented or top-down method, which is often used in product design. Winning, extraction and processing of metals are part of the interconnected worldwide metals production system linked to product design, manufacture and recycling. This complexity cannot be dealt with using LCA. This paper will discuss a dynamic model that deals with this complexity, at the same time providing detailed bottom-up insight into the complex web of materials referred to as Metal Ecology. An example on the replacement on lead in solders will be discussed to illustrate this.

Jan 1, 2005

By M. Modis

According to previously established theoretical analysis and under certain conditions, a critical sampling grid can be determined for an earth-related space-distributed natural variable. Sampling above this critical limit adds little to the mapping results, while based on this limit the ideal process of reproducing the original phenomenon is theoretically defined. The objective of this paper is the application of the above theory to the Klidi lignite deposit in Florina, Greece, where Public Power Corporation (Greece) has conducted a sampling campaign to estimate the deposit thickness and its heating value. Data acquired for these two variables were analysed in the space domain in order to derive their respective correlation functions. The structural analysis of the deposit generated two covariance models, one for the thickness and one for the heating value, from which the critical sampling grid was identified. The conclusion drawn is that in some parts of the deposit the sampling grid is denser than required. Finally, an optimal sampling grid is proposed in order to gain the maximum information at the lowest cost.

Jan 1, 2006

By R. Edward Williams

The Spa1lation Process Because of the low thermal conductivity of many hard rocks, rapid heating of these rocks produces a thin surface layer in which the temperatures attain high values. Thermal expansion of this surface layer is constrained by the reminder of the still cool rock, and when stresses within the surface rock become high enough, the surface rock breaks away from the cooler rock behind it and flies or falls off as a thin flake called a spall. Then the next, newly exposed surface is heated, and the process continues. This process is the basis of spallation drilling. The hot gases from a jet burner provide the heat for spallation to occur, and their high velocity provides a scouring action that transfers heat to the rock and removes the spalls as rapidly as they form. Spallation is a process which works in very hard rock. It is dependent upon the thermal expansion coefficient and the thermal diffusivity of the rock but is also affected by any discontinuities in the rock. To date the efforts which have been made to evaluate the various rock according to their spallability has been minimal. As the success of this process is dependent upon the characteristics of rock it is expected that the study of rock mechanics will prove to be of greater value to this program than to the other mechanism for drilling and excavating rock. Commercial Uses of SPALLATION In the 19408s, the Linde Air Products Division of Union Carbide (UC 1 began developing spallation for use in mining taconite ore, which is presently the chief source of iron in the United States. In this work UC developed a jet-piercing tool that burned fuel oil with oxygen to produce spallation and contained mechanical cutters to remove rock that was not amenable to spallation. The UC jet-piercing machines have since produced about 40 million feet of shallow blast holes used for emplacing explosives in the taconite mines. During this work it was found that hole diameters could be increased by merely reducing the advance rate of the burners and that existing holes could be enlarged by making another pass through the hole with the same burner. The Browning Engineering Go. of Hanover, N.H., has developed a hand-held spallation burner to cut slots in granite. It has been used for a quarter of a century and is now standard equipment for quarrying granite throughout the world. This burner, which resembles a small jet engine oriented with its exhaust pointed downward, is the forerunner of a flame jet burner used to spall experimental holes in granite at maximum rates in excess of 100 ft/hr when operating in hard, competent granite. It uses No. 2 fuel oil, which is burned with compressed air. The system uses water to cool the burner and the exhaust gases. These gases, along with the steam produced from the cooling water, blow the spalls from the hole. Experimental Work Theoretical and experimental work has been accomplished by the Massachusetts Institute of Technology and the Los Alamos National Laboratory. This work is reported in Refs. (3) and (4). To verify the experimental results of this work laboratory scaled down field tests were conducted using two we1 1 characterized granites from quarries in Barre, VT and Westerby, RI, under defined heating conditions. In the laboratory tests a propane - oxygen heating torch was used to direct a flame at the granite surface and the spal 1 ing process was examined at various heating rates with a high-speed video taping system operated at 200 frame per second. This produced a time-lapse sequence where the onset of the spallation process was easily distinguished. Also the heat flux from the torch to a flat surface at various stand off distances and flows was measured. A similar set of tests was conducted using the more easily quantified and uniform heat source of a 1.5 kw GO2 laser. This allowed accurate

Jan 1, 1986

By James S. Browning

This article describes research conducted by the University of Alabama's Mineral Resources Institute on removing sulfur from coal using a combination of gravity concentration and flotation methods. The research used the gravity and centrifugal force capabilities of Humphrey's spiral to remove part of the pyritic sulfur and ash. Flotation of the Humphrey's spiral coal concentrate removed additional sulfur and ash. In addition to obtaining good recoveries of high quality coal, the tests demonstrated that pyrite in the wastes could be isolated as enriched flotation products by the combined gravity-flotation approach. Domestic coal production in 1980 reached a record 757 Mt (835 million st). Forecasts are that coal production and use will approach 907 Mt (1 billion st) by 1985, with much of the increased production being used for electrical generation. Of current production, it is estimated that about 454 Mt (500 million st) of coal nationwide are subjected to some degree of cleaning, ranging from crushing and screening to gravity concentration and flotation. With expanded coal production and more rigid controls placed on coal combustion, it is reasonable to assume that the tonnage subjected to some form of preparation will increase, with time, thereby adding to the burden of coal rejects being generated. A major problem associated with increased coal use for electricity generation is the need for utilities to conform to emerging federal air quality standards for sulfur oxide emissions from power plants. To a lesser extent, other industrial operations burning coal face a similar situation. Controlling sulfur oxide emissions by using low-sulfur coal or resorting to stack gas desulfurization are two recognized approaches, but removing sulfur from the coal, if possible, by physical beneficiation before combustion is preferred because of its simplicity and lower cost. Reducing the inorganic sulfur in coal to tolerable limits by physical beneficiation, including flotation, as now practiced, is still relatively inefficient and has not kept pace with the more sophisticated physical beneficiation approaches generally employed by the nonfuel minerals industry. As one phase of a broad program conducted by the Mineral Resources Institute at the University of Alabama on sampling and evaluating impounded fine size waste coal, tests were made using samples from two impoundments to determine the quality of coal recoverable by a combination of Humphrey's spiral concentration and flotation. Particular attention was given to organic and inorganic sulfur deportment and distribution in the various products. Another objective was to determine if pyrite in the waste could be recovered for possible marketing. The use of Humphrey's spirals to concentrate fine coal was practiced as early as 1947. At that time, the Hudson Coal Co. used spirals to concentrate anthracite fines. In 1950, spirals were used to treat fine-size bituminous coal and later, in 1976, spirals were reportedly used for removing pyrite from fine-size coal. More recent investigations have discussed the use of Humphrey's spirals for treating fine-size waste coal in Alabama. Samples Description Washer waste samples were obtained from two impoundments. One sample was representative of waste generated during washing steam coal from the American, Pratt, and Nickel plate seams in Alabama. The other sample represented metallurgical coal washer waste from processing coal from the Carter, Johnson, and Milldale seams. Partial samples analyses are shown in Table 1, and screen analyses of the samples are given in Tables 2 and 3. Screen ana¬lyses showed that one sample contained about 4% of material coarser than 2 mm (10 mesh) and the other contained about 13% of plus 2 mm (10 mesh) material. Also, the analyses indicated that ash and sulfur increased as the material's particle size decreased. Experimental Laboratory Tests Sink and float tests using a heavy liquid (1.4 specific gravity) were made on the different screen sizes shown in Tables 2 and 3 to determine the effectiveness of gravity concentration for separating coal from pyrite and other gangue constituents. Tests indicated that gravity separation, under optimum conditions, could be expected to yield coal products containing 10-12% ash, and 0.2-0.3% pyritic sulfur. Good coal recoveries could also be expected. Using sink-float tests as a guide, tests were made in a Humphrey's spiral by a procedure in which a 90-135 kg (200-300 lb) portion of the

Jan 8, 1982

By Jonathan Ludlow, Robert A. Jankowksi

Introduction A longwall operator can make few changes to increase output, significantly reduce respirable dust, and decrease power consumption. Reducing drum speed, and thereby cutting with increased pick penetration, is one. This article defines the benefits of deep cutting in terms of reduced dust production and power consumption. It also identifies the practical aspects of high pick penetration in terms of shearer performance and coal loading. Before examining some practical aspects of reducing drum speed and looking at the theoretical background, it is worthwhile to summarize what is meant by high penetration and deep cutting, and what potential benefits and pitfalls may be expected. Deep cutting (in the sense of high penetration rather than wide web) can be defined in one or more of the following ways: • Cutting with an average pick penetration distance higher than that used in the past. • Cutting with a pick penetration higher than the longwall operator would have used if the advantages of deep and slow cutting were not considered. • Cutting with a well-designed shearer drum below 40 rpm. All these definitions are slightly arbitrary. They are given to provide a basis for discussion and to make the point that any move towards deeper, more efficient cutting can result in operational benefits. The benefits of deep cutting appear in many different areas. The most noticeable benefit, provided suitable instruments are available, is the reduction of airborne respirable dust. During an experiment on a longwall in the Pittsburgh seam, a nearly four to one reduction in dust levels was seen when drum speed was halved. Not all studies have shown such a big reduction, but it seems that some benefit is almost always obtained when drum speed is reduced. Production rate and specific power consumption are also affected (in a positive sense) by reducing drum speed or increasing pick penetration. Although these changes may not be as spectacular as those in dust level, they contribute to the economic return of the longwall operation. Similarly, improved washability through fines reduction may have a beneficial economic effect. Cutting with shearer drums operating at lower speeds does have some possible deleterious impacts that an operator should be aware of. For example, cutting reactions - loads imposed on the picks by the coal being cut - will be increased as a deeper cut is used. Steps must, therefore, be taken to ensure the stability of the shearer and provide an adequate haulage effort. These increased cutting reactions also result in higher loads on the power transmission system (gearboxes, ranging arms, pick boxes, etc.) from the shearer motor(s) to the pick tip. These higher loads must be anticipated and provided for with the necessary hardware. In particular, extra haulage power must be provided with low drum speeds, since haulage effort required increases roughly in proportion with pick penetration. Because the drum will be rotating more slowly or will have fewer picks, the load on shearer components will also be more variable. If suitable, robust equipment is not used, this increased vibration will decrease reliability. Benefits of Deep Cutting Lower dust levels, decreased specific power consumption, and improved product washability are the most noticeable benefits of reduced drum speeds. Although the benefits will vary greatly with mining conditions and the type of coal, some examples of what can be expected are described below. Reduced Dust Levels Figure 1 shows principal results of a study on the effects of reduced drum speed conducted on a longwall in the Pittsburgh seam (Ludlow, 1981). This figure shows that average dust production was reduced by about 70% when drum speed was halved. By making some assumptions about such quantities as coal density, it is possible to apply this proportional reduction to the quantity of respirable dust liberated per ton of coal mined. When this is done, two kinds of results are obtained: • At 70 rpm, about 1 g (15 gr) of airborne respirable dust is created for every ton mined (roughly one part per million). At 35 rpm, only 0.28-0.37 g/t (3.9-5.1 gr per st) of coal mined become airborne respirable dust. • At 35 rpm, nearly four times the amount of coal may be mined before the compliance level is exceeded, compared with 70 rpm.

Jan 3, 1984

By N. Djordjevic

Introduction The vibration energy that travels beyond the zone of rock breakage is wasted - all it does is cause damage and annoyance. Under favorable geologic conditions, this energy may travel many kilometers before it finally drops bellow the background noise level. This has forced the introduction of more restrictive limits to the allowable vibration levels. To maintain the profitability of a mine or quarry and to comply with environmental protection legislation, blasting needs to be performed in a more predictable way that will ensure that the final results are optimal in both mining and environmental terms. Minimization of ground and induced structural vibration is, therefore, a matter of good blasting practice -not just an attempt to avoid complaints from neighbors. A new approach to blast vibration minimization is presented. The method utilizes the existence of the local minimums in the frequency spectrum of single-hole blast vibration. The results are presented as a delay map showing maximum ground or structural vibration velocity for many combinations of interhole and interrow delays. A blast engineer can use this map to select an optimal combination of delays that will produce a low vibration velocity at a selected location and, at the same time, will be acceptable from the point of view fragmentation or muck pile shape. Response of structures A typical single-story building has two distinctive forms of vibration. The first form is a translatory vibration of the building superstructure (frame), and the second form is a flexural-bending vibration of the structural elements such as walls and floors (Siskind et al., 1980). Walls and floors will have several modes of oscillation. The most important is the fundamental mode of the vibration, as this mode will have the lowest frequency, the largest amplitude and will cause the largest deformations. The vibration of walls and floors is also a source of secondary vibration within a building, due to the movement of loose objects, e.g., furniture and wall fittings (Dowding, 1985). The subjective estimate of the amplitude of a vibration by the occupants of a house will be unrealistically high due to the effect of the secondary vibrations. It is now well recognized that human perception of ground vibration begins at levels that are far below those that will produce minor damage to the most fragile structures. In mining and quarrying operations, complaints regarding vibration levels tend to be determined more by human perception or annoyance rather than by observations of damage. Ground vibrations on the order of 0.08 mm/s (0.0032 ips) (Walter and Walter, 1979; Murray, 1987) can be felt, and they are often subjectively estimated to be about 100 times greater. Amplification of ground vibration in the structure depends on the amount of energy in the ground vibration spectrum that is in the vicinity of the structure's resonant frequencies together with the damping ratio of the structure at these particular frequencies. One way to avoid significant vibration amplification in a structure is to shift the dominant part of the ground vibration away from the resonant frequency range of the structure. The presented method allows for the selection of the optimum blast timing (interhole and interrow delay times) to minimize the resultant ground vibration amplitudes in a given frequency range. It also allows the selection of suboptimal, but still acceptable, blast delay times, which may be preferable from the point of view of overall blast performance. Determining frequency ranges of importance in structures In attempting to avoid amplification of vibrations in the structure, ground vibration amplitude of frequencies close to the resonant frequencies of the building needs to be minimized. The width of the frequency-minimization region is determined by the value of damping in the building. If the damping ratio is low, then the frequency-minimization region around the resonant frequency will be relatively broad. By shifting the dominant part of the ground-vibration energy into the frequency region much higher than the resonant frequency of the building (1.5 to 2 times), it is possible to achieve efficient vibration isolation of the building. In such cases, the structural dynamic amplification of ground vibration will be less than unity. The first step in this minimization process is the determination of the structural resonant frequency. From the frequency spectrum, it is possible to determine the resonant frequency of the structure. Sometimes the

Jan 1, 1998

By W. J. Schlitt

Introduction A major uncertainty in assessing the potential for heap-and dump-leach projects is how to determine the extraction-rate curve for the recovery of the mineral values from coarse ore. Such material could either be run-of-mine (ROM) or primary crushed ore. The problem with field testing coarse ore, especially for new projects, is the large scale and extended leach times needed to accurately determine the final extraction-rate curve. At least 5 x 103 to 5 x 104 t of representative ROM ore are typically required for a copper test heap, and much more is often used. Kennecott, for example, recently constructed a 0.9 Mt (1 million st) ROM test heap at the Bingham Canyon Mine in Utah. In such coarse ore operations, the ultimate level of extraction will require a leach cycle that can extend from several months to a few years. Quite often, project development schedules do not provide the luxury of mining such large quantities of material or running such long tests. Instead, test data are usually limited to results from column leach studies on relatively fine ore, often with a top size that does not exceed 25 mm. Maximum leach times are also short, typically less than a year before an initial decision is needed on project viability. Proposed method One approach to estimating the recovery from a coarse ore leach is to assume that the leach solution will have some ultimate penetration distance into the rock. Then, the final level of mineral extraction in this "leached rim" will equal the ultimate level of extraction identified in various testing programs. Obviously, if the radius of a given rock fragment is less than the penetration distance, that fragment will be fully leached at the end of the operation. With larger rock sizes, the percent recovery will fall off as the size increases and the fraction of unpenetrated rock mass increases. Such an approach sounds simple but is likely to be complex when applied to a real project. For example, the penetration distance will be a function of both the rock characteristics and the effective length of the leach cycle. The important rock characteristics include rock porosity, the degree of internal fracturing and the mode of mineral occurrence. With regard to the latter, penetration is likely to be greater if the leachable mineralization occurs on fracture surfaces or in veinlets, as opposed to fine grains uniformly disseminated throughout the rock mass. An estimate of penetration distance may be derived from column or heap tests by noting the depth of solution penetration into the larger rock fragments after three, six and 12 months of leaching. While the penetration rate is ore specific, something on the order of 10 to 20 mm/y may be appropriate for competent, primary copper (chalcopyrite) ore. For gold in tight quartz, the rate may be about the same. Copper oxide ores and gold that is hosted in a more porous rock matrix are likely to have penetration rates that are at least two to three times higher, and an even higher rate should be appropriate for uranium hosted in sandstone. As noted above, the length of the effective leach cycle is likely to be measured in years. On this basis, the ultimate penetration distance (dp) would vary from less than 50 to several 100 mm when a particular ore is leached to exhaustion. Several sets of mathematical manipulations are necessary to convert a rock size distribution and corresponding value of dp into an estimated extraction-rate curve. The first step is to break the ROM size distribution down into intervals and then calculate the radius for the mean rock size in each interval. This is shown in Table 1 for rock sizes up to 1.75 m (about 6 ft) in diameter. The next step is to calculate the volume of unleached core and the fraction of rock that is leached. This is done for the following three values of dp: 25, 100 and 250 mm. Results are shown in Table 2. The third step is to select the ultimate level of recovery that will be achieved in the fraction of material that is effectively leached, i.e., the outer zone that is penetrated by the leach solution. This is clearly a site-specific factor that can only come from metallurgical test results on representative ore

Jan 1, 1998

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

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

Jan 1, 1994

By L. C. Lang

INTRODUCTION The introduction of 165-mm (61/z-in.) holes to underground mining operations has made possible the application of Canadian Industries Ltd's (CIL) vertical crater retreat (VCR) mining method, illustrated in the accompanying sketches. This unique and revolutionary new application of spherical charge technology (see the Appendix), when applied to primary stopes and pillar recovery, eliminates raise boring, slot cutting, and dilu¬tion of ore by backfill; greatly improves fragmentation; reduces labor and time requirements; eliminates uphole drilling and blasting; and minimizes or completely elimi¬nates damages by blasts to the walls and retreating backs of the stope or pillar. If vertical large diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at the calculated optimum distance from the back of the undercut and detonated, a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, min¬ing of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed. The VCR method is also applicable to drop raises and has the potential to replace all other raising methods under most circumstances. PILLAR RECOVERY Levack Mine Inco Metals Co., Ontario Div., provided the first opportunity for the method in pillar recovery. The Levack mine's high grade ore pillar No. 4800 on the 975-m (3200-ft) level was used for the production¬scale experiment (Figs. 1 to 3). The pillar was about 49 m (160 ft) long, 6 m (20 ft) wide, and 20 to 26 m (65 to 85 ft) high. The mined area on both sides of the pillar was backfilled with a 1:30 cement:sand mixture. The pillar was removed in two phases. In phase 1, the standard uphole method was used to blast down the 18-m (60-ft) long section of waste from the bottom of the ore into the undercut. From the pillar's top sill, 165-mm (61/a-in.) holes were drilled downward into the pillar, and by measuring the depth of the holes, the results of the uphole blast were determined and roof line 1 was established. The bottom of each hole was plugged, then filled with sand to place the center of gravity of each spherical charge (loaded from the top sill) at a predetermined optimum distance from the horizontal free surface. The charges were then detonated. After detonation, both draw drifts were full of extremely well-broken material. The depth of each hole was measured again, and the plot of these depths re¬sulted in roof line 2. The same blasting procedure was repeated and the resultant new back elevation was marked by roof line 3. The poor results of the initial uphole blast at one location (notice the peak in area 1) appeared to influ¬ence the subsequent new backs. A third blast success¬fully evened the back, and resulted in roof line 4. An unblasted slab averaging 6.3 m (20.9 ft) thick remained below the pillar's top elevation as the final sill. In all three spherical charge blasts fragmentation of the blasted material was extremely fine. The backfill was fully exposed on both sides of the now-blasted pillar. The backfill remained undamaged and the ore was not diluted by sand. The remnants of all the 165-mm (61/2 -in.) holes remained clean and undamaged, and the holes had well-defined bottoms that could be easily measured and plugged. Each blast took down a 3.9 to 4.2-m (13 to 14-ft) thick horizontal slab of ore. Productivity was three times greater than that of the previously practiced cut-and-fill method. Since this was the first such experiment, blasting the remaining 6-m (20-ft) thick final slab was the sub¬ject of some deliberation. If the described method was repeated, we could have ended up with a 1.8-m (6-ft) thick sill unsafe to work on. It was therefore decided to blast the whole sill using two spherical charges prop¬erly placed in each hole, but with the application of

Jan 1, 1982