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By Dermot M. Ross-Brown

Introduction Slope angles are one of the major factors affecting the shape of the final pit and the location of the walls Due to differences in geology, the optimum slope angles vary from pit to pit and within different parts of the same pit. Indeed, it is not unusual to have different slope angles for the top and bottom portions of the same cross section. The optimization of these slope angles is a major study in itself and this chapter is intended as a general introduction to the subject. It is not intended to be a comprehensive text for the absolute beginner The objective is to put the various aspects of slope stability analysis into relative perspective by describing the vari¬ous analytical techniques available and indicating when each may be most appropriately used References are made to the many excellent texts and papers on slope stability for the reader who does not have a formal training in soil or rock mechanics. In writing this chapter, an assumption has been made that a tentative mining plan has been drawn up, includ¬ing a preliminary estimate of the slope angles. It has also been assumed that all the necessary field and laboratory data has been collected and reduced to a form suitable for input into the stability analyses.

Jan 1, 1979

By D. W. Boydell

The response of reclaimed gold tailings to acid leaching was investigated in the laboratory, followed by a pilot-plant program in which the operations of CCD, CIX and SX were tested with plant feed material. These results formed the basis for the design of the extension to the existing uranium plant at Blyvooruitzicht Gold Mine, commissioned in July 1977. This circuit consists of slimes-dam reclamation by bucket-wheel excavator at a rate of 100,000 t solids per month, acid leaching in pachucas, a five-stage CCD washing section, two parallel C1X absorption columns and two elution columns, SX and an ADU precipitation section. Barren solution is recycled to the CCD section as wash and to the repulper at the slimes dam. The performance of the plant during the first year of operation is compared with the predictions on which the plant design was based. Rated capacity was achieved within three weeks of start-up at a nominal dissolved uranium recovery in the ion-exchange section of around 98%. Owing to recycle, the over-all recovery in the ion-exchange and solvent extraction sections is somewhat higher than this figure.

Jan 1, 1979

By Robert F. Hauck

This chapter discusses a new technology for computer control in open-pit mine operations: the real time automatic dispatching of manually driven trucks to shovels. The monitoring of truck activity is accomplished by detection of local radio signals. Computer-operated directives are communicated to drivers by remote controlled digital display units inside the truck cab. The production objective is to maximize tonnage in a cost effective manner. The problem is formulated in the appendix and the solution method is briefly described. Side conditions such as maintaining delivered ore grade aims, balancing waste to ore volumes to keep within stripping ratio tolerances, and meeting production rate requirements temporarily established for certain shovels are also discussed. Experience with a developed event- to-event simulator applied to analyses of several North American mining operations using this technique is summarized. Complete systems employing this technology are now under- going installation at mine sites in Canada and the United States.

Jan 1, 1979

By Ivan V. Klumpar

If components of a gas-liquid stream react with each other and with stationary solids, involving parallel and sequential, reversible and irreversible, and slow and instantaneous reactions, the process can be defined by a set of simultaneous differential rate equations and simulated as multistage semibatch re¬actors. Each stage is divided into finite time increments. Within each increment, even the simultaneous reactions are treated as sequential. The proposed method follows process evolution from non-steady state startup to normal operating conditions. It allows computing byproduct buildup in recycles, tracer response curves and, in case of underground processes, expansion to adjacent parts of the deposit.

Jan 1, 1979

By Charles J. Kucera

Today's highly publicized issues of resource development and environmental protection are not new. But recently these issues have been played one against the other, and some would have us believe they are mutually exclusive. Why? Because modern man has realized that the shortest route to achieving economic progress is to subdue nature. He extracts mineral resources from the earth to provide the raw materials upon which other processes and services are based. But often in his search for economic progress he finds that other highly- held values (e.g., an unspoiled environment) have been violated. The natural response is a counter action - to attempt to protect nature. But this response is basically unsatisfactory because it is reactive.

Jan 1, 1979

By D. Raicevic

After more than 93% of the uranium is extracted from Elliot Lake uranium ores by a sulphuric acid leaching process, the leach residue (tailings) contains small amounts of uranium and radioactive isotopes, particularly radium-226, which is the most serious health hazard. Heavy metal components and pyrite are also present, along with the gangue minerals. Currently, over /000 acres of the Elliot Lake area are covered with these tailings, which contain over five million tons of pyrite. Because of constant oxidation of the pyrite by bacteria (thiobacillus) and the presence of moisture, pyrite slowly generates sulphuric acid, which steadily leaches the metal-bearing constituents from the tailings. The seepage-flows of the dissolved contaminants, although often quite low in volume, have an environmental impact on the Elliot Lake area. Although these seepages are treated and most of the contaminants removed and impounded, a small amount of the radioactive contaminants reaches Lake Huron via the Serpent River. This paper describes a flotation approach for treatment of the Elliot Lake uranium tailings to produce new, decontaminated tailings practically free of pyrite, with radium, thorium and uranium contents considerably reduced. The decontaminated tailing produced, which comprises about 75% by weight of the current uranium tailings, appears to be suitable for mine backfill. Because mine backfilling normally uses about 50% of the plant tailings, the surface storage of about half of the uranium tailings, therefore, would be eliminated by this process. Mine backfilling would also increase the mine production and thus enlarge the over-all uranium resources due to recovery of the ore from pillars. The pyrite concentrate produced from the current uranium tailings would be suitable for sulphuric acid production. The possibility of uranium recovery, and disposal of radium and thorium from the concentrates produced, is now being studied at CANMET.

Jan 1, 1979

By Robert M. Dreyer

Although lateritic ore deposits are relatively flat-lying, surficial ore bodies, the evaluation of lateritic bauxite and nickel deposits is among the most difficult problems confronting an economic geologist. Laterite is a soil formed under tropical conditions resulting in the concentration, within the soil zone, of certain trivalent metals - principally iron and aluminum, but also, locally, nickel and manganese. The soil zone is divided into two parts: (a) the true soil or lateritic zone and (b) the zone of partially altered bed rock or saprolite. Whereas bauxite generally is confined to the lateritic zone, some nickel deposits occur in both the laterite and saprolite zones. The evaluation of bauxite and lateritic nickel deposits thus has both similarities and differences. These ores have been formed by supergene, soil-forming solutions moving downward through zones of widely varying permeability within the weathering rock. Because the permeability varies widely, both vertically and laterally. the extent of weathering may vary greatly within short distances. Because the weathering process varies widely, the degree of metal concentration may also vary widely. It is not at all uncommon for a laterite to change, within a few centimeters, from a virtually valueless rock to a high grade ore. Moreover, the depth of weathering, and hence the depth and form of the ore profile, vary greatly. Obviously, for this type of deposit, a few grab samples or a few drill holes are virtually worthless. The evaluation of such a deposit requires a sophisticated statistical study based on thousands of carefully selected samples. As a rule of thumb, the cost of evaluating a large bauxite or lateritic nickel deposit will be about one cent per ton of ore in the ground. Thus, the cost of evaluating a several hundred million ton deposit may be considerable. The following factors must be considered during the evaluation: Type of Sampling: The best and most reliable samples for the evaluation of lateritic ores are taken along large, vertical channels in hand-dug pits. Where labor costs are high, some form of drilling may be used in place of pitting, but it is always wise to choose a drilling procedure after checking drilling results against a series of pit channel samples. The drilling may not reproduce exactly the results from pit channel samples, but it may be possible to derive a constant correction factor. If the physical character of the laterite varies within the deposit, it may, of course, be necessary to change the type of drilling, the correction factor or both. In some areas, it has been found empirically that augers, chum drills, vacuum drills, rotary drills, or core drills yield satisfactory results. The problem is to find which drilling technique gives results most similar to channel samples through the heterogeneous sequence characterizing the laterite in a given area. In lateritic nickel deposits, the drilling characteristics of the lateritic and saprolitic portions of the deposit are very different. Within the saprolite, the major nickel values are commonly found in weathered zones within less altered boulders - a type of occurrence which makes sampling very difficult.

Jan 1, 1979

By R. S. Norman

The objective of this program was to select approximately 100 acoustic materials for flammability evaluation. Noise control programs have been initiated on several types of mining equipment and vehicles which can be used in underground coal mines. Often, the materials which are used in these noise reduction programs have not been tested for flammability properties according to the same reference standard. During this program a variety of materials were selected for measurement according to ASTM E-162 procedures. A radiant heat energy source was used to measure flammability properties of 15.2 x 45.7 cm. samples. In the absorptive materials category, 45 percent of the materials tested had a flame spread index of less than 25. For barrier, damping, and isolation materials the corresponding values were 40, 43, and 9 percent.

Jan 1, 1979

By Frank W. Schmechel

The subsidence monitoring program at the Old Ben No. 24 Mine was planned in 1976 as part of the on- going demonstration of longwall mining. The construction and installation of the network of surface subsidence monuments was started in Play of 1976 and completed four weeks later after which an initial reference survey was made of all surface monuments. Mining of the first longwall panel started in September of 1976 and was completed in May of 1977. Mining of the second panel started in August of 1977 and is continuing at this time. The Automatic Data Acquisition System (ADAS), which is the subject of this paper, was installed over the second panel on March 29, 1978 near the panel center (Figure 1) and approximately 104 meters (370 feet) in front of the face position at that time. Mine No. 24 is located approximately 5 kilometers (3 miles) Northwest of Benton, Franklin County, Illinois. Coal is being mined from the Illinois (Herrin) No. 6 seam which is at a depth of about 189 meters (620 feet) at the site. The Herrin No. 6 is up to 2.7 meters (9 feet) thick in the mine but a constant thickness of 2.1 meters (7 feet) is being mined by the demonstration longwall with both roof and floor coal being left for stability. Although the seam is flat laying at the mine, it dips north- ward regionally. Topographically, the surface over the demonstration panels is essentially flat and low laying with several areas of standing water. The Rend Lake Dam (United States Army Corps of Engineers) is about 305 meters (1000 feet) north of the demonstration site.

Jan 1, 1979

By R. J. Davidson, M. V. Nkosi, V. Veronsee

When activated carbon is used for the recovery of gold and silver from Witwatersrand gold-plant solutions, severe fouling of the carbon by calcium carbonate is experienced. Also, difficulty is encountered in the retention of the adsorbed silver on the carbon because it is displaced by gold. The acidification of plant pregnant solutions before they are contacted with carbon was therefore investigated as a means of overcoming these shortcomings. The results from batch laboratory experiments on the effect of pH on gold adsorption from plant solutions and from synthetic solutions of high ionic strength indicated reaction rate optima in the region of pH 5 and below pH I. The effect of pH on the equilibrium capacity constant for gold adsorption onto activated carbon was also studied, using both a plant solution and a synthetic solution of high ionic strength. In the case of the plant solution, the results suggested only a marginal change in carbon capacity for gold in the pH range 12 to 6, with significant increases in capacity at lower pH values. In the case of the synthetic solution of high ionic strength, the capacity constant was found to increase steadily as the pH was lowered from 12 to I. It was also found that the gold adsorption capacity of the carbon was significantly better for the synthetic solution. No precipitation of gold from the pregnant solution (gold content approximately 6 g/t) was noted as the pH was decreased to values as low as ,. However, about 25 per cent of the 0,6 g/t silver content was found to have precipitated when a pH value of 3,0 was attained. Further substantial precipitation of silver took place as the pH was lowered further. The investigation was extended by the passing of suitably acidified gold-plant solutions through one or more columns containing activated carbon. Preliminary results showed that the particle size of the carbon had a significant effect on the rate of gold adsorption. No displacement of adsorbed silver by gold was observed to take place at an influent pH of 5. In a typical continuous 'merry-go-round' operation involving three columns in which gold was eluted from the lead column every 48 hours, effluent values averaging 0,001 g/t gold could be consistently achieved under simulated plant-operating conditions. The results were obtained at relatively high velocities of influent solution, resulting in very little hold-up of gold in the carbon-recovery circuit. More than 99,9 per cent of the gold and more than 99,8 per cent of the silver was recovered in the adsorption circuit, while the elution data indicated an average elution recovery of 99,8 per cent of the adsorbed gold and 98,2 per cent of the adsorbed silver. Elution of the carbon was carried out at 91 QC. This involved pretreatment of the carbon with a sodium cyanide solution, followed by elution with deionized water. At a loading flowrate of 66 bed volumes per hour, the indicated take-up of metals on the carbon in the lead column was as follows; 53 k g/t gold, 6 k g/t silver, 57 k g/t copper, 14 k g/t nickel, and I k g/t zinc. Only a trace of cobalt was adsorbed. At an elution rate of 1,7 bed volumes per hour, the lead column was found to be completely stripped in under 5 hours, with concentrations of gold in the eluate reaching 6 to 7 g/l. Besides the excellent adsorption and elution results obtained when acidified pregnant solution was used, no fouling of the carbon was indicated, and the pressure drop across the three columns was considered to be well within practical limits. When compared with conventional zinc precipitation, the carbon process would appear to have considerable economic merit, with strong indications that significant savings would be possible both in capital and operating costs. Further pilot-plant experiments, preferably on a gold plant, would be required before a meaningful cost comparison could be made with the existing process.

Jan 1, 1979

By R. Samuels, R. Klimpel

"IntroductionLaboratory and industrial grinding tests have shown that the process of size reduction of metal containing ores can be significantly influenced by chemicals added to the slurry being ground.,The terms grinding aid or grinding additive refer to a substance which, when mixed into the mill contents, causes an increase in the rate of size reduction. The increased rate can be used to grind a higher feed rate to the desired product size or it can be used to produce a finer product size at a fixed rate. Whether the use of a grinding aid is justified in any given situation depends on the cost of the substance versus the improvement of output or product quality obtained with its use.A wide variety of grinding additive mechanisms have been suggested in the general mining literature (1,2,3,4,5,6,7). The prevention of particle agglomeration, slurry cushioning of grinding media impact, grinding media coating, changing the strength of macro porous rocks, and chemical adsorption on flaw surfaces to change bonding forces and surface energy have all been suggested. More recently, some grinding additive work has been carried out (8,9,10,11) which suggests that the use of efficient chemicals to control the fluidity of dense slurries is a viable economic approach to achieve improved grinding rates. These increased grinding rates are achieved with no increase in power to the grinding mills and apparently with no increase in media consumption costs. Unlike much of the previous work with grinding additives, this recent work is accompanied by a rather complete engineering characterization and prediction of performance capability. The chemicals involved were chosen so as to efficiently control the viscosities of slurries under conditions of high shear yet not adversely affecting flotation, flocculation, and other downstream surface chemistry operations."

Jan 1, 1979

By R. J. Munday

Although uranium deposits in northern Saskatchewan are clustered around the Athabasca basin, basement lithostructure determines their distribution. Basal Wollaston Group pelites are invariably graphitic and carry uraninite in felsic segregations and biotitic pegmatite. Where these intersect the Athabasca Formation, remobilized uranium concentrated as pitchblende. Uranium is also concentrated in silicified and hematized sandstone where the Athabasca Formation overlies post-orogenic faults. Presumably, pregnant solutions associated with the fault were stripped of their metals as they emerged into the contrasting physico-chemical environment of the sandstone basin. Both these types of mineralization correlate with a Grenvillian (-1,000 m.y.) heating event and associated diabase intrusions. However, the rich pitchblende veins at Uranium City appear to predate the Athabasca Formation, but display spatial correlation with the Martin Formation. They are tentatively associated with a syn- to post-D3 retrograde metamorphic event which thoroughly leached crystalline basement and generated uraniferous fluids which escaped, and precipitated metals, into fractures associated with faults that define the grabens in which the Martin Formation was deposited. Remobilization events allowed some of this uranium to move upward into fractures within Martin sediments.

Jan 1, 1979

By Aurel Goodwin

Remarks of Aurel Goodwin, Chief, Division of Health, Metal and Nonmetal Mine Safety and Health, Mine Safety and fEealth Administration, U. S. Department of Labor, before the Society of Mining Engineers of the AIME, Littleton, Colorado, on October 18, 1979. I appreciate this opportunity to speak before the Society of Mining Engineers of the AIME. The topic of my presentation is the Mine Safety and Health Administrations's (MSHA) concern and obligations with regard to asbestiform minerals. Asbestiform minerals - the magical group of hydrated, silicate minerals which possess a fiber like structure capable of being woven, like and into cloth. Their unique combination of properties, such as resistance to heat and chemical attack, high tensile strength, and flexibility was recognized centuries before Christ. Pausanias, the early Greek geographer, speaks of golden lamps made about 430 B.C. with incombustible wicks of "Carpathian flax." The Romans used asbestos as cremation clothes to conserve the ashes of deceased persons of rank. Charlemagne had an asbestos table cloth that he used to pass through fire to clean (1). Early recognition of the health hazards associated with exposure to asbestos was documented in the first century by Pliny the Elder, a Roman naturalist, and Strabo, a Greek geographer. Bath wrote of slaves with sickness of the lungs, whose occupation was that of weaving asbestos into cloth. This concern for the health of asbestos workers disappeared over the ages.

Jan 1, 1979

By Harrison R. Cooper

Approximately 40 concentrators throughout the world have successfully adopted computer control for flotation circuit operation since 1969. Successful use of flotation control, gaining up to 2% efficiencies, has largely derived from effective employment of on-stream slurry analyzers and integration of key measurement data through digital processors. Requirements for reliable measurement have also led to improved functional reliability in magnetic flow meters, pH, reagent feed, and other key elements. Methods used in computer control of flotation will be reviewed and variations in operating philosophies will be described with reference to the approaches taken to flotation automation.

Jan 1, 1979

By Alden B. Carpenter

Oil-field brines have been collected and analyzed from Jurassic and Cretaceous rocks in the Mississippi Salt Dome Basin, from Jurassic and Lower Cretaceous rocks, southwestern Arkansas, and from Ordovician and Silurian rocks, southern Michigan. All of these brines are genetically related to evaporites and appear to be brines expelled from halite-bearing rocks, presumably during compaction. The brines in rocks closest to the halite-bearing rocks are uniformly low in lead and zinc suggesting that the metals in metal-rich brines are not acquired within evaporite sequences. The metals in metal- rich brines in Mississippi and in Russia appear to be derived from red beds and may have been originally associated with ferric iron oxyhydroxides. Reduction of these compounds by migrating hydrocarbons appears to release lead and zinc to the associated brines accompanied by the fomation of authigenic ankerite.

Jan 1, 1979

By S. J. Bluhm

Recommendations are given for the design of multi-stage spray chambers in which air flows horizontally through water sprays. A detailed description is given of one such spray chamber at Hartebeestfontein gold mine that is used for bulk cooling of the incoming ventilation air, and that has been constructed according to these recommendations. A method for predicting the performance parameters of horizontal multi-stage spray chambers is given, and the method is illustrated with some examples. The method is equally applicable to spray chambers in which the water, rather than the air, is cooled.

Jan 1, 1979

By Rao P. K, Sahoo P. K

The main constituents of ferromanganese slags are silicates of Mn, Ca and AI. Factors affecting the recovery of manganese from a low manganese content ferromanganese slag (about 16 per cent MnO) during acid leaching were studied. These included time, temperature, agitation, sulphuric acid addition, solid:liquid ratio and particle size. Agitation rates greater than 300 rev/min. leaching periods greater than 30 min and temperatures above ambient were found to have no effect on Mn recovery. Dissolution of Mn was found to be highly dependent on the amount of acid added.About 90 per cent of the Mn was recovered when 5·5 times the stoichiometric amount of acid was used. About 28 per cent of the Mn was recovered when the stoichiometric quantity of acid (with respect to Mn alone) was used. However, the recovery increased to about 75 per cent when the total stoichiometric requirement of acid with respect to Ca, Al, Mg and Fe as well as Mn was used for leaching. The extent of Mn dissolution also depended on the particle size of the slag. The percentage of Mn recovery (using 5·5 times the stoichiometric amount of acid) increased from about 60 per cent to about 92 per cent when particle size decreased from -36 +52 to -150 +200 mesh (B.S.S.). When the results of this study were compared with those in a previous study involv,ing the leaching of a high manganese content slag (containing about 40 per cent MnO) results were similar except that there was a gr-eater consumption of acid in the present case (i.e. low manganese content slag).

Jan 1, 1979

By Annie G. Smelley

The Bureau of Mines, U.S. Department of the Interior, as part of its mission to effect pollution abatement, conducted research to devise means for disposing of phosphatic clay wastes and reclaiming mined land. Part of the investigation was performed in a cooperative research project with 10 phosphate companies in Florida. The colloidal properties of the phosphatic clay suspensions, called slimes, make them difficult to dewater and dispose of without impoundment storage aboveground. Various reagents to flocculate and dewater these plant wastes were evaluated, and a polyethylene oxide polymer was found to function in a specialized flocculation dewatering technique. The flocculation dewatering method was investigated in bench-scale batch tests and applied in continuous dewatering tests using standard mineral-processing equipment. In continuous tests at feed rates of up to 58 liters per minute, wastes at 4 percent solids were dewatered to about 20 percent solids. About 0.8 kilogram of polymer was used for each metric ton of dry waste clay solids; processing time was about 4 minutes. Requirements for the treatment system are (1) proper mixing of the clay-waste feed and flocculant, (2) rapid removal of water as it is released, and (3) physical manipulation to promote cohesion of solids.

Jan 1, 1979

Provide reasonably priced, reliable fire protection for blasthole drills. Approach An automatic fire sensing and suppression system using both dry and gaseous chemicals was developed, and is more shock and vibration resistant than similar industrial fire protection systems. How It Works Temperature sensors mounted in the drill's machinery house, hydraulic room, operator cab, and transformer area detect the heat from a fire, and trigger a sequence of actions. The same fire fighting sequence can be initiated from manual alarm boxes in the cab, hydraulic room, or machinery house. Electrical power to the drill Is immediately cut off, and a horn sounds to warn the drilling crew. After a 40-second, built-in, delay to allow time for drill's ventilating fan to stop, gaseous and dry chemical fire extinguishing agents are piped to and released in various rooms of the drill. The gaseous extinguishing agent (Halon 1301) is released in the machinery house, hydraulic room, and operator cab where dry chemical clean-up from electrical panels and machinery would be difficult and costly. Dry chemical is piped to the transformer room at the rear of the machine. Location of the principal components of the system are shown in the drawing.

Jan 1, 1979

By C. R. Pelley

This paper describes the analytical characteristics of Western subbituminous coal and the manner in which these characteristics apply to their use in a utility-sized steam generator. It compares these characteristics with those of typical Midwestern coal and Eastern bituminous coal. The paper also reviews in detail the design considerations that must receive attention when various blends of these three coals, or any combinations thereof, are planned to be used. This research paper describes variations in analytical and utilization characteristics when coals from different origins are blended in order to achieve a predetermined sulfur content for the purpose of predicting sulfur oxides emissions as they affect environmental regulations and the resulting alterations in utilization characteristics such blending would cause, and how such blends of coal would affect a utility-sized steam generator.

Jan 1, 1979