Search Documents

By Ann M. Hagni

The mineralogical, chemical, microtextural, and particle size character of electric arc furnace (EAF) dusts are being studied by reflected light microscopy, scanning electron microscope-energy dispersive spectroscopic analysis (SEM-EDS) and electron probe microanalysis (EPMA). The dusts consist predominantly of metallic iron, iron oxide, and calcium-iron-manganese silicate phases, together with smaller amounts of iron-manganese-zinc-calcium, iron-calcium-zinc-manganese, and calcium-iron-manganese silicates. The samples vary in their zinc contents. Chlorine occurs in phases containing iron, lead, zinc, manganese, and calcium. The dust particles range greatly in size from 1 to more than 300 ~m, but most particles are about 1 to 6 ~m. Larger particles. about 20-50 um across. also are present. They are predominantly spherical in shape, but some are cenospheres, and others are more angular particles, some of which are broken spheres and cenospheres. Their internal microtextures exhibit evidence of liquid immiscibility, exsolution, skeletal crystal growth, and oxidation react i on rims. Because EAF dusts have been classified by EPA as hazardous wastes, there is heightened interest in the treatment of these dusts to render them non-hazardous. The mineralogy of products of pyrometallurgical treatment of EAF dusts by flash furnace and pyrohydrolysis are currently under study.

Jan 1, 1991

By A. Kevin Beaton, Robert W. Laakso

The desire for dry passageways promotes the NATM as it provides a final support system most suitable for the installation of a waterproofing membrane. Waterproofing during construction is a cost effective way to enhance safety and function as well as increasing the life of an underground structure. The NATM approach utilizes the surrounding mass as an integral element of the "ringlike" support system. At the Pennsylvania Turnpike Commission Lehigh Tunnel No. 2 various rock classifications were determined to require support measures ranging from 2 to 12 inches of shotcrete along with spiling, wire mesh, deformed bar, lattice girders and rock bolts as needed. Field instrumentation and monitoring programmes are a key factor for the success of the NATM. When the support "shell" stabilizes smoothing shotcrete, waterproofing and a final liner can be placed.

Jan 1, 1991

By R. L. Brown

The Buchans mining district is located roughly at the center of the island of Newfoundland and contains some 18 separate mineral deposits. The first of these deposits was discovered in 1905, the latest discovery having been made in 1979. To the present (1983), some 16 Mt (17.7 million st) grading 1.33% Cu, 7.5% Pb, 14.5% Zn, 115 g/ t (3.7 oz per ton) Ag, and 1.4 g/t (0.04 oz per ton) Au have been mined. Mining operations are now largely terminated, but a few hundred thousand tons of economically marginal reserves may be mined in a salvage operation. The Geological Association of Canada Special Paper No. 22, "The Buchans Orebodies, 50 years of Mining and Exploration," contains papers which describe the history, geology, and the application of evolving geological concepts in exploration. Papers by Neary, Swanson, and Thurlow and Swanson in that volume are topical to the present discussion. Successful exploration techniques at Buchans included physical prospecting, utilization of the state-of-the-art geophysical techniques, saturation diamond drilling which was made possible by the high gross value of the existing ore, and, since 1955, diamond drilling directed by fairly well understood geological parameters. The record of discovery started in 1905 when the Buchans River deposit was found by Matty Mitchell, a prospector employed by the Anglo- Newfoundland Development Co., a British company which owned mineral rights over a very large area in the Red Indian Lake-Exploits River drainage system.

Jan 1, 1991

By C. A. Summers

The U.S. Bureau of Mines applied process mineralogy to investigate a lead and zinc deposit located near Story Creek, Colville Mining District, Alaska. Mineralization is hosted by the Kayak Shale in veins, veinlets, and breccia. Sphalerite and galena, with supergene anglesite, occur in symmetrical quartz veins and between breccia clasts. The study sample analyzed 2.11 percent Pb, 5.37 percent Zn, 0.014 percent Cu, and 1.2 oz/ton Ag. Analyses of crushed products showed that 33 percent of the sample weight, and more than 50 percent of the Pb and 65 percent of the Zn, are in the minus 850 um fraction. Liberated grains of sphalerite and galena were found in crushed products finer than 300 um. Beneficiation tests indicate the material is amenable to processing by traditional gravity and flotation techniques.

Jan 1, 1991

By R. H. King

The US coal industry has been using continuous room-and-pillar mining (CRPM) machines for more than 30 years for longwall development. This equipment has also been used for production mining in areas that are not appropriate for longwall mining. From 1977 to 1987, the portion of annual US production mined by continuous miner (CM) sections grew slightly from 65% to 70%. Over this time, mine operators and manufacturers have made some advances to attain the full production potential of this method. This is evidenced by the fact that average tons per eight hour shift for continuous miners in the US grew from 180 t (200 st) in 1977 to 431 t (475 st) in 1987. The authors (1990) reported that the increase in productivity was due to several factors: • a drop in prices; • a more mature workforce; • union stability and incentive programs; •procedural improvements; and •technological advances. The authors estimated that about one-third of the increased production came from the first three and other unaccounted for factors and one-third each for the last two. Even though improvements were considerable, continuous miner capacity is much higher than 431 t/shift (475 st per shift). For example, the machinery is capable of mining more than 9 t (10 st) per machine for the 6.5 available hours per shift or 3.5 kt (3900 st). However, this has never been attained on development. The averages of most mines are less than 900 t/shift

Jan 1, 1991

Optimisation of capital in the development process is critical to the profitability of any new mining venture. The process of optimisation should apply to all aspects, that is the financing, management, revenue, reporting and, of course, the development of the project. Unfortunately, too much emphasis has been placed on the optimisation of the financing or reporting aspects and too little on the project itself. This paper addresses the latter.

Jan 1, 1991

By A. Masutomi

Although great strides have been made during the past two decades in the reduction of the severity of accidents and injuries in underground coal mining, the number of accidents resulting in disabling injuries and lost time has continued to increase. Looking ahead to the 1990s, the reduction of accidents and incidents typically found in these latter categories should be emphasized in safety training for underground coal mining. Loss control is an approach which shows promise for applications in mine safety training since loss-control strategies focus on the pre-contact stages of an incident rather than on the incident, itself. Applied to safety training, the heightened awareness of personnel to workplace hazards at the pre-contact stages should result in a reduction of accidents and injuries. This paper explains the concept of loss control and shows how it can be applied in the development of the needs assessment for miner training programs. An example, based on research conducted at the continuous-miner worksite, is provided.

Jan 1, 1991

By K. G. (Ken) Thomas

"This paper gives an insight into the various processes available to treat gold ores. A brief overview is given of the various types of refractory gold ores. Particular emphasis is given /0 comparing the costs of oxide vs refractory gold ore processes, inclusive of grinding through to dore bar.HistoryThro ugh time gold has been recovered from the earth's crust by various mean s. Five major historical steps can be identified for metallurgical processing as shown in Table I .Gravity concent ration started with, for example, river bed panning and developed to more sophisticated processes such as the Johnson concentrator, belt concentrator and jigs. The South African gold fields were a major developer of gravity concentration processes during the first half of the twentieth century.As coarse free gold became depleted, stamp milling, in conjunction with amalgamat ion , was developed. Basically, rock was crushed under large mechanical hammers an d the gold collected in mercury . The mercury was then evaporated in a retort to leave sponge gold for further fire refining.As deposits became deeper and gold finer, the gold became more difficult to recover ; fine grinding was required to liberate the gold. In the 18905, MacArthur and the Forrest Brothers of Glasgow, Scotland, introduced the cyanide pro cess for dissolution of gold and its subsequent precipitaiton by means of zinc. This process, in vogue until the early 1970s, was commonly referred to as the Merrill-Crowe process for recovering fine gold .Activated charcoal was used as a means of recovering gold from solution in the 1880s, especially in chlorination plants in Australia( l). Two factors probably retarded the development of carbon instead of zinc-based processes during those year s. First , there was no satisfactory procedure for the elution of carbon to allow its recycling and, second, the development of the Merrill-Crewe process for gold recovery by zinc cementation probably discouraged any further work on the use of carbon"

Jan 1, 1991

By A. W. Bryson, J. D. le Roux, B. D. Young

Several rate models for the adsorption of gold cyanide from aqueous solution onto activated carbon are given in the literature. The purpose of this study is to critically examine their ability to simulate the adsorption process over a wide range of conditions. An account is given of a comprehensive set of experimental work on the kinetics of gold adsorption onto a coconut shell- based activated carbon. Equilibrium isotherms were also determined. It is concluded that the homogeneous surface-diffusion model fits the experimental data reasonably well and provides a good balance between simplicity and descriptive accuracy. The model contains two parameters: the extra-particle mass-transfer coefficient and an effective intraparticle surface-diffusion coefficient. Both can be evaluated readily from simple adsorption experiments.

Jan 1, 1991

By U. Lauwers

The influence of different waste materials in sinter blends on the quality and mineralogical composition of a sinter is described. As a result of partial preservation of the original textures of waste materials in the final product the degree of assimilation of the original material in the sinter product can be ascertained. Recycling of waste materials such as dolo char , arc-furnace slag, scrap fmes, mill scale and BOF-mud via the sintering process a saving on costs of raw materials such as flux and fuels, and a reduction on pollution can be achieved. A comparison of a typical sinter with a sinter produced through the utilization of waste materials showed a distinct correlation between certain types of waste material and some metallurgical properties. Dolochar in a sinter blend acts as part of the fuel and flux that are normally added to a sinter mix. Generally dolochar improves the quality of the sinter and decreases the consumption of coke. Other waste materials adversely influence the quality of the sinter. The glass and calcium ferrite content of a sinter when dolochar is used in the blend is similar to that of a sinter produced with only coke as a fuel in the sinter mix. The glass content is lower and the calcium ferrite content is higher, in a sinter produced from dolo char in the blend compared to a sinter produced with additions of other waste materials. Changes in the mineralogical composition are related directly to the metallurgical properties of the sinter product.

Jan 1, 1991

By Pat DeWilliam

The Yakobi nickel-copper deposit is located essentially at tidewater in the east-central part of Yakobi Island, about 16 1 airline km (100 airline miles) west of Juneau, and can be reached by float plane and boat from Juneau and Sitka. The fishing village of Pelican, population 400, is situated 14.5 km (9 miles) to the east. An unimproved road, approximately 4 km (2.5 miles) long, connects a dock site at Lisianski Strait with the main part of the deposit known as the Basin Area (Fig. I). Other mineral occurrences are known as the Takanis, Flapjack, and Beach Areas. Another smaller nickel-copper deposit is located about 32 km (20 miles) to the south on Western Chichagof Island and is known as the Mirror Harbor nickel-copper deposit. The original discovery of the Yakobi deposit was made by S.H.P. Vevelstad about 1920 and a 47.5-m (156-ft) long adit was driven on the southwest side of the deposit in the Basin Area. The deposit was examined by J.R. Van Fleet of Union Carbide in 1930 and by J.G. Fairchild in September 1931. Stanley and Yates of the International Nickel Co. of Canada (Inco) visited the property in August or September 1937. In 1942, the controlling interest in the deposit, covered by 124 unpatented claims, was said to be held by Vevelstad and associates

Jan 1, 1991

By S. Banisi, J. Marois, A. R. Laplante

"This paper compares the behaviour of gold and ore in the grinding circuit of Hemlo Gold Ltd. Particular attention is given to grinding kinetics, classification behaviour, and liberation. For grinding kinetics, laboratory grinding of gold flakes and silica, 840-1200 um in size (14/20 mesh), was also performed to provide a standard for comparison and to study in a controlled environment the mechanisms of breakage, folding, smearing and impacting.Grinding kinetics were characterized at laboratory scale with breakage and selection functions. The breakage function showed that 75% of gold disappearing from the 840-1200 um class reported to the 600-840 um; this figure is much higher than for silica ( 46%) or most other minerals, and implies that comparatively fewer fine fragments are produced from the breakage of gold flakes, on account of gold's malleability. The selection function of silica was four times that of gold. At plant scale, the ratio of the selection function of ore and gold increased from 6 for at 50-100 um to 20 at 500-1000 um.Observation of original flakes and their fragments with the scanning electron microscope showed that the main mechanism of ""grinding"" from the 840-1200 um to the 600-840 um class was folding rather than actual breakage. This was confirmed by a measure of individual flake weight. The postulated breakage cycle is that relatively globular flakes become increasingly serrated through grinding; they can then regain their original shape through folding and reflattening into cylindrical and spherical shapes, as the majority does; however, some fragments will at this point undergo actual breakage. Photographic evidence is presented to support the proposed mechanisms.Smearing of gold unto silica was observed when silica was ground after a gold, and embedding of silica into gold flakes when gold was ground after a silica. Both mechanisms occur whenever gold and harder minerals are jointly ground, and result in ""unliberation"" of gold. It is postulated that smearing of gold onto other minerals plays an important role in explaining how gravity recovery can increase overall gold metallurgy, especially when used ahead of flotation.Classification of gold in the primary hydrocyclones takes place at a much finer cut-size than ore, 20 vs. 57 um. This and the slower grinding kinetics result in high gold circulating loads in the primary loop, 6700%, and a gold content in the primary cyclone underflow (PCUF) is twenty four times that of the ore. The gold size distribution for primary cyclone overflow is much finer than that of the ore, although they are essentially similar in the PCUF.Liberation, or more accurately ""gravity recoverability"", was assessed with a 7.5 cm (3"") Knelson concentrator. It varied between the thirties for the primary mill discharge and two cyclone overflows, to highs of 89 to 92% for the PCUF and secondary mill discharge, 89 and 92%, respectfully."

Jan 1, 1991

By M. W. Grabinsky, M. S. Diederichs, E. Hoek

Provided that they are used correctly, numerical models can be of considerable assistance to mining engineers in designing underground excavations. Guidance is given to the engineer or geologist who is not a specialist in numerical analysis but who is interested in using numerical modelling tools or in employing someone to analyse the stresses, displacements and failure in the rock surrounding underground excavations. A description of the theoretical background to a number of these numerical models and discussion of the advantages and disadvantages of each are followed by guidelines on the selection of appropriate models for each stage in the development of a mine, from exploration to the completion of mining

Jan 1, 1991

By Thomas M. Barczak

Wood cribs are used extensively to stabilize openings in underground coal mines. This U.S. Bureau of Mines report describes the results of full-scale testing of several multitimbered wood crib configurations. Multitimbered configurations are wood cribs that are constructed with two or more timbers per layer. Nine different crib configurations were evaluated. Configuration parameters under study included the aspect ratio (height-to-width relationship) and the percentage of solid rr:aterial used in crib construction. Test results and generalized wood crib behavior are discussed and recommendations are made concerning crib constructions to maximize strength and stability. Crib performance is enhanced by reducing buckling effects by decreasing the aspect ratio and geometry considerations that increase the moment of inertia. A cost benefit analysis of multitimbered wood crib configurations is made. Generally, increasing the percentage of solid provides more resistance at less unit construction cost. A comparison of laboratory results and underground observations of several wood pack wall supports suggest in situ load capacity is less than that measured in laboratory testing. Reasons for the apparent decrease of in situ load-bearing capability include load rate effects, stiffness of reaction surfaces, and varying boundary conditions that alter the load transfer mechanics of the crib structure.

Jan 1, 1991

By Alan J. Reed

The Afton copper-gold mine is located in the southern interior of British Columbia, 16 km (10 miles) west of Kamloops and 257 km (160 miles) northeast of Vancouver. 'An aerial view of the property in 1976 is shown in Fig. 1. The Afton copper-gold ore body [27 Mt (30 million tons)] of open pit ore with a grade of 1.0% copper, 0.46 g/t (0.01 6 oz per ton) gold, and 4.1 g/t (0.12 oz per ton) silver, occurs within the Triassic Iron Mask pluton adjacent to its faulted contact with Tertiary rocks. The geological setting of the Afton ore body is described by Carr and Reed (1976). Rocks of the Paleozoic, Mesozoic, and Cenozoic eras outcrop in the vicinity of Kamloops, but the unit with the greatest economic significance is the Upper Triassic Nicola volcanic group and its subvolcanic dioritic intrusion, the Iron Mask pluton. The Afton ore body occurs at the northwesternextremity of the Iron Mask pluton immediately south of a deep graben structure that developed in post-Triassic time (Coast Mountains orogeny) and which was the focus of vigorous sedimentation and volcansim in Tertiary times. The Iron Mask pluton is a lenticular composite intrusion approximately 14.5 km (9 miles) in length in a northwesterly direction. The petrography of the Iron Mask pluton ranges from picrite and serpentinite to latite porphyry. Outcrops of individual petrographic units are generally elongate in a northwesterly direction parallel to the axis of the pluton. The Iron Mask pluton represents the moderately slowly cooled core of an oceanic fissure volcano which built up an island flanked by shallow-water sediments including reef knolls and which, during its later stages, intruded its own volcanic pile.

Jan 1, 1991

By David G. Chedgy

Noise control is an important consideration in the design and operation of coal preparation plants. It is essential to understand and control noise sources to provide a safe environment for operating personnel. The average noise exposure levels in a typical coal preparation plant generally range between 95 and 100 db. An operator should not be exposed to noise levels of this magnitude for more than 4 and 2 hr, respectively, according to MSHA (Mine Safety and Health Administration) regulations. The major noise sources in a coal preparation plant include noise from impacting the inside surfaces of steel chutes, vibrating screens, coal crushers, pumps, motors, fans, and blowers. The magnitude of noise varies considerably from one situation to another. NOISE EXPOSURE IN COAL PREPARATION PLANTS Noise regulations established by the Occupational Safety and Health administration (OSHA) are accepted by MSHA relative to caal preparation plants. Fig. 14-43 shows a graph of the time of noise exposure in hours per day to which a worker can be exposed for each A-weighted sound level. According to this graph, a worker is permitted to spend 8 hr in a 90-db noise environment, but that time must be halved when the A-weighted sound level increases 5 db. The noise dosage a worker receives is determined by the ratio of the length of time the worker spends in a particular noise environment divided by the length of time he is permitted to spend in that environment. If the worker is exposed to several different sound levels, his total dosage would be the sum of each of the individual dosages. The equation for determining the dosage is: where C is the actual duration of exposure at a given steady state noise level and T is the noise exposure limit for the level present during the time C According to MSHA regulations, the dosage, D, should not exceed one (1) for any worker for a full day of work. As an example of typical plant noise levels, vibratory screens and crushers emanate noise levels in excess of 90 db at a distance of 3 ft (0.9 m) with no material flow. With material flow, noise level increases 1 to 10 db depending

Jan 1, 1991

By I. J. Brown

Paper prepared for presentation at Industrial minerals and the environment, held in Leeds, 24-35 September 1990. The need to submit an environmental statement (EIS) has certainly increased the workload of those concerned with both its production and assessment. However, there is now good evidence that the system is enhancing public confidence in the minerals industry in particular and improving relationships between the developer, planning authority and those consulted such as statutory bodies, parish councils, civic societies, nature conservation interests and ramblers

Jan 1, 1991

By G. B. Amicucci

Golden Sunlight Mines Inc. operates an open-pit gold mine and cyanidation plant about 11km (7 miles) northeast of Whitehall, MT. Golden Sunlight is a subsidiary of Placer Dome US Inc. Mining is accomplished with conventional open pit drilling, blasting and hauling methods. The mining rate is about 102 kt/d (113,000 stpd). Gold occurs as fine-grained disseminations in a pipe-shaped breccia and in immediately adjacent country rock. The ore body is low grade, averaging 2 g/t (0.059 oz per st) of gold. The metallurgical plant uses conventional carbon-in-pulp (CIP) technology. Daily throughput averages 6.5 kt (7000 st). Recoveries are enhanced through a sand tailing retreatment facility while a preaeration process reduces operating costs. The low grade ore dictates that maximum mill throughput at the desired grind size be maintained. Therefore, in October 1988, an Armco Autometrics (now Denver Autometrics) particle size monitor (PSM) 400 was commissioned. The PSM 400 is an on-line instrument. It measures particle size and percent solids in a slurry stream. The PSM pulls its own sample, deareates the sample and passes it through diametrically opposed ultrasonic sensors. The unit uses ultrasonic sound to measure the deareated slurry stream's particle size and percent solids.

Jan 1, 1991

By R. P. Killmeyer, R. E. Hucko, P. S. Jacobsen

Cleaning of three coals from the eastern US was tested by five selective agglomeration processes and one selective flocculation process. The coals and processes are described, and the results are reported and compared. Two of the processes, which ground the coal to the finest volume mean diameter, met the stringent test program quality specifications.

Jan 1, 1991

By R. L. Attridge

Namibia's newest mining enterprise, the Navachab Gold Mine, was officially opened by President Sam Nujomo during June 1990. The mine is a joint venture between Erongo Mining & Exploration Company Limited, Metal Mining Corporation of Canada, and Rand Mines Windhoek (Pty) Ltd. Erongo has a 70 per cent interest in the mine, while Metal Mining and Rand Mines hold 20 per cent and 10 per cent respectively. Anglo American Corporation and its associates hold two-thirds of Erongo shares and CDM (Pty) Ltd one-third. The Mine is managed by Erongo Mining & Exploration Company. The property is located some 10 km southwest of the town of Karibib and 170 km west-northwest of the Namibian capital of Windhoek (Fig. 1), the name of the Mine being taken from the farm on which the gold deposits were discovered. Navachab is an opencast operation and was brought into production in December 1989 at a cost of around R84 milion-RS,2 million under budget. The plant is designed to treat 840 kt annually at an average gold grade of 2,6 g/t. This wil produce 1900 kg of gold per year, which at today's prices will earn Namibia more than R60 million in foreign exchange.

Jan 1, 1991