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By D Craw

Bulk ore tailings from the Macraes gold mine gold extraction process are stored in a large (ca 0.6 km2, 80 m deep) mixed tailings impoundment whose pH is maintained near 8 by natural rock neutralizing capacity. The mixed tailings will be stored in this impoundment in perpetuity after mining ceases. Confidence in the long term environmental stability of these tailings can be gained from examination of mineralogically and chemically similar geological situations in the immediate vicinity. A sequence, typically about 5 m thick, of sands and gravels derived from the Macraes mineralized zone 12 000 to 28 000 years ago contains rounded detrital sulphide mineral grains which are unoxidized despite close proximity to the surface and occasional incursion of oxygenated waters. Saturating water pH is currently 7-8. Sands with 0.2 to 0.8 wt % organic carbon host sulphate-reducing bacteria (SRB), and localized cementation by authigenic framboidal pyrite has occurred. Comparison of the 28 000 year-old sediments with the modem tailings suggests that similar chemistry will prevail, possibly with deposition of authigenic pyrite in the tailings, in the long term. As long as the present slightly anoxic and circumneutral pH environmental conditions are maintained in the mixed tailings dam, sulphide decomposition and acidification are unlikely. These sediments provide a plausible geological analogue for long-term benign storage of the mixed tailings impoundment.

Jan 1, 2000

By Yoshinori Iwasaki, Shosei Serata, Yutaka Mochida

A new comprehensive insitu instrumentation technique named this RCC Method has been developed for rapid assessment and improvement of the structural stability of underground openings. When applied to tunnel construction, the degree of present and prospective safety of underground openings can be determined in quantitative terms within 10 to 20 minutes of microcreep observation. When applied to mining, roofs and floors can be stabilized by controlling the stress conditions. This stress control method has eliminated the need for intensive roof bolting and extensive floor renewal in all cases where it was applied. Use of this technique is illustrated by field examples from the Seikan Undersea Tunnel of Japan and deep potash mining in Canada. NEW INSTRUMENTATION TECHNIQUE: THE RCC METHOD Failure problems in underground excavation usually arise in weak non-elastic grounds where conventional rock mechanics theories of elasticity are not only useless but also misleading. Presently, discontinuity planes in the ground are being considered a factor in assessing these failure problems, and while this development is interesting academically, it is not of much use in actual excavation work, where it is nearly impossible to identify and quantize all existing planes of discontinuity. Even if such an extraordinary task could be accomplished at an underground excavation site, effective use of the data would be difficult and impractical for rapid excavation work, with the possible exception of rare cases of ideally elastic ground.

Jan 1, 1976

By A. Sinha, N. K. Babar, H. R. Kalihari, A. K. Chakraborty, M. R. Saharan, C. P. N. Pathak

A case study is presented on the optimization of cable bolting patterns in cut-and-fill stopes at the Chikla Mine, MOIL, India. Rock mass characterization, using Bieniawski’s RMR and Barton’s Q, indicated that the stopes lie in the “no supports required” to “spot bolting required” support categories. In contrast, the current practice of cable bolting in a 2.0 m by 2.0 m grid pattern for stope widths less than 12 m displays different results. Numerical modelling results are in line with the rock mass characterization study. It can be affirmed, as well, that no cable is carrying rock loads greater than 5 t. This fact is derived from observational behaviour of the crown pillar of an experimental stope with open stoping operations and hanging cables in the cut-and-fill stopes of the mine. An increase in the cable bolting grid spacing is, hence, recommended for implementation by mine management, based on the study described in subsequent sections. It is also suggested that the mine adopt the open stoping method, provided results of a rock mechanics instrumentation program in an experimental stope with a cable bolting grid of 2.5 m square are favourable.

Jan 1, 2007

By J. S. Harington

Our purpose is to record and to explain the widely changing numbers of the huge total of black workers on the gold mines of South Africa over the century from 1896 to 1996. These men were recruited from up to twenty territories, and their contributions are examined. The major areas providing labour over this long period have been Lesotho, the former Transkei within South Africa and Mozambique. From 1989 to 1996, for instance, these three areas accounted for over sixty per cent of the total labour force (tlf). Our principal Figure 1 shows the tlf fluctuating through time to reach a peak of 480 000 in 1988. Irregular falls are also documented with the most serious in 1995 to 180 000. Based on this graph we consider the political, economic, sociological and historical reasons behind these changing employment opportunities. These have included external world-scale events such as war or depression, as well as internal changes of national economy and policy. We show the importance for numbers employed of the discovery of gold in the Orange Free State in 1946, the influence generally of the ?Winds of Change? and the firm maintenance by the State of law and order. Varying gold production costs and profitability against high tax rates, allied with increasing wages and a lack of development capital, have led in the last decade to employment falling by fifty per cent, despite an actually higher value of gold output. Finally, this analysis considers the role of the migrant labour system as it affects, and is affected by, conditions in the gold mining industry, with special reference to its less than humane consequences in the numerically dominant territories of Lesotho and Transkei.

Jan 1, 2004

A GREAT deal of attention has lately been given to the discussion as to the education required for engineers, and especially for mining engineers. The Institute of Civil Engineers has had a committee's report on it which has come to some interesting and valuable conclusions.The Institute of Mining and Metallurgy had a few years back a very long discussion on the subject, and the subject appears to be one of perennial interest to similar societies. Two recent addresses by Prof. Kernot, as president of the Victorian Instituteof Civil Engineers, are among the most interesting and practical of the contributions to the discussion. Engineers seem to feel either that the coming generation has not the same capacity to benefit by the training that has made themselves what they are, or that such training has turned out in themselves inefficient men, and that they must consequently struggle to obtain a better one for their successors. For not only is there a desire to extend existing instruction to cover the advances of science, but there seems to be a general desire also to make a completely radical change in the methods.It is not necessary to observe that education does not mean the imparting and acquiring of information, but rather the cultivation of the faculties so that the best use can be made of them and of any information acquired. For instance, a man is not educated who has by heart his lectures and the contents of his text-books, but that man is who remembers little of the actual wording of his text-books, but is able to find anything required in them...

Jan 1, 1909

By L. Flett, S. Woods, C. Deredin

"Following a significant reduction in world Imperial Smelting Furnace (ISF) capacity from 2003 to 2008, a major optimization initiative was successfully carried out on the lead upgrading and bulk (zinclead) flotation circuits in the Brunswick Concentrator.This paper describes the process improvements that have taken place in order to reduce the quantity of bulk concentrate:•,Increase in lead upgrading rougher flotation circuit capacity•,Installation of a closed-circuit bulk regrind utilizing MillPebs and 150 mm Cavex hydrocyclones•,Installation of two additional cleaner flotation stages on the bulk circuit•,Optimization of the manipulated variables in the lead upgrading and bulk circuitsINTRODUCTIONThe Brunswick Mine (Xstrata Zinc Canada) is located near Bathurst, New Brunswick and is one of the world’s largest lead/zinc mines. Milling of ore commenced in 1964 from the No. 12 orebody, with an expansion in 1966 to accommodate ore from the No. 6 open pit mine. The planned closure of the mine will take place at the end of the first quarter in 2013. The ore is a complex massive sulphide, consisting of approximately 20% valuable minerals, 60% pyrite and 20% silicates.A simplified flowsheet for the concentrator is given in Figure 1. The overall flowsheet involves SAG milling, secondary ball milling, selective flotation, and dewatering by thickening, filtering and drying. The complexity of the flowsheet evolves from the requirements for fine grinding and the production of four concentrates: zinc (sphalerite), lead (galena), copper (chalcopyrite) and bulk (sphalerite plus galena).Primary hydrocyclone overflow feeds the copper-lead aerator-roughers for flotation at pH 9.0 with soda ash. Plant xanthate (80-20 sodium isopropyl and potassium amyl) and AP407 are used as collectors in the copper-lead circuit. The rougher concentrate is directed to the copper-lead regrind circuit and the tailings are fed to the secondary grinding circuit. The secondary hydrocyclone overflow from the secondary grinding circuit feeds the copper-lead rougher-scavenger banks. Rougher concentrates are reground to a P80 of 28 µm prior to three stages of copper-lead cleaning. The first stage of cleaning is done in an open circuit with the tail reporting to the cleaner scavenger. The cleaner scavenger concentrate reports to the regrind circuit and the cleaner scavenger tail combines with the rougher/scavenger tail to feed the zinc circuit. The second and third stages of cleaning are performed in closed circuit with the cleaner tail being recycled to the copper-lead regrind circuit. The average final concentrate grades 23% Pb, 3% Cu and 704 g/t Ag, and is fed to the copper separation circuit."

Jan 1, 2013

By A. D. Thomas

Some years ago, the authors developed an analysis of non-Newtonian pipe flow based on enhanced turbulent microscale and associated sublayer thickening. This analysis, which can be used for various rheologic modelling equations, has in the past been applied to smooth-walled pipes. The present paper extends the analysis to rough walls, using the popular Bingham plastic equation as a representative rheological behaviour in order to prepare friction-factor plots for comparison with measured results. The conditions for laminar-turbulent transition for both smooth and rough walls are also considered, and comparisons are made with earlier correlations and with experimental data.

Jan 1, 2007

By A. S. Moghaddam

The transportation of cemented and non-cemented mine tailings in the form of paste through pipelines is one of the main stages of paste backfill operations. For the purpose of pipeline design, the rheological properties of the materials, such as yield stress and viscosity must be determined. In this study, the yield stress of two Canadian mine tailings, for gold and copper mineralization, were studied using the vane method and slump tests. The comparison between the vane method and the slump tests indicates that the slump tests generally overestimate the yield stress values for both cemented and non-cemented paste backfills. In addition, the effects of 5% Portland cement (type 10) and mixing time on the yield stress of paste backfills were investigated. The percentage of cement as a binder agent has a direct effect on the increase of the yield stress of paste backfills. Yield stress measurements show that mixing times of 4 minutes and 12 minutes have the same effect on the yield stress values.

May 1, 2007

By Claude Bazin, François Robichaud

"The control of the operation of semi autogeneous grinding (SAG) mills is mainly achieved through adjustments of feed rate, water addition and speed of rotation. With the introduction of image based sensors it is possible to have on-line measurements of the size distribution of the ore feeding a SAG mill. This information could be used for the development of a control strategy to regulate a SAG mill feed size distribution to a set-point. The control of the size distribution is carried out by adjustment of the speed of the ore extractors underneath the ore stockpile. The potential and difficulty associated with the implementation of such control are assessed using a multivariable statistical analysis of data collected from the operation of the grinding circuit of the Agnico-Eagle Goldex mine in Val d’Or, QC. A control strategy for the ore feed size distribution is assessed using a simulator of the ore stockpile operation.INTRODUCTIONThe popularity of autogeneous (FAG) and semi autogeneous (SAG) grinding mills has been increasing over the last 20 years (Napier Munn et al., 1996). These mills use the ore or a combination of ore and steel balls as the grinding media. Since the ore is used as a media for the size reduction the mill ore feed size distribution impacts the performance of the mill. Several papers were recently published on this subject (Dance et al.,2011; Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011) showing the advantages of preparing the feed size distribution to maximize the circuit throughput or mill efficiency (Wennen & Murr, 2011; Dance & Valery, 2011; Allaire et al., 2011). With the availability of particle size sensor based on image analysis of the ore on a conveyor (Partanen et al., 2001; Bobo et al. 2004; Edwards et al., 2002; Wei & Craig, 2009) it is possible to control the size distribution of the ore feeding autogeneous or semi autogeneous mills (Partanen et al., 2001).This paper describes a method to control a size distribution index to a set point via the manipulation of the speed of the extractors that withdraw the ore from the stockpile feeding the grinding circuit. The paper consists of 4 sections. The first section describes the considered grinding circuit. A statistical analysis of the collected data on the grinding circuit is presented in the second section. This analysis demonstrates the necessity of using real-time adaptation mechanisms to develop a control strategy for the size distribution of a grinding circuit. The third section describes the method used to simulate the ore stockpile operation and the last section describes the preliminary development and assessment of a control strategy of size distribution of the ore feeding a SAG mill."

Jan 1, 2013

By Narasimhan K. S, Sastri S. R

A typical high silica low grade magnesite from the Salem area of Southern India, was used in studies to develop a suitable flotation system involving commercially inexpensive sodium oleate as the collector together with other reagents which can give a product of the required grade with an acceptable recovery. It was found that the presence of metallic ions such as Cu+, Fe+++ or Ca++ introduced by the addition of their salts, or by their formation with the addition of hydrochloric acid to react with magnesium and calcium carbonate in the me, can act as effective depressants for silica. The addition of 'hydrochloric acid' was found to give best results.

Jan 1, 1980

By T. Belem

The main objective of this study is to better understand the effect of underground CPB placement conditions on its microstructural, hydro-mechanical and geochemical properties. To reach this objective, three PVC columns, each 3 m high, were built, and were filled with CPB material. The columns allow simulating undrained (UD), half-drained (HD) and fully-drained (FD) conditions along with the measurement of resulting self-weight consolidation settlement of the CPB. Bleed and drainage waters were also collected from each column for geochemical analyses (pH, Eh, EC, SO42- content). After 90-days of curing time, a total of sixty CPB samples were cored from the backfill columns and subjected to uniaxial compression tests followed by physical characterization (Gs, void ratio, water content, degree of saturation, specific surface area). The results showed that the total drainage water and the maximum observed settlement occur mainly within the first 72 hours after the columns are filled. The corresponding volumetric strains suggest that in situ backfilled stopes simulated behave in a similar way to the fully-drained (FD) or the half-drained (HD) conditions. The MIP test results show that the pore structure varies between the top and the bottom of each column.

May 1, 2007

By Michael W. Neill, O&apos

Drilling slurry has become an effective aid to the construction of drilled shafts in caving and/or waterbearing soils. Historically, drilling slurry made of bentonitic clay and water was first used in the US in water well drilling as early as 1889. In coastal Texas drilled shaft contractors began using slurry, often made of on-site clay mixed with potable water, in the years following World War II, on a relatively routine basis. This type of slurry was used to "process" boreholes, in which the borehole was drilled "in the dry" until water or unstable soil was encountered, after which water and on-site clay were introduced into the borehole and mixed in place with the auger or other drilling tool. Drilling then continued under a slurry consisting of the uncontrolled mixture of clay and water until the excavation had reached a relatively impermeable and cohesive material. At this point, a steel casing was dropped into the hole and sealed into the lower stratum of impermeable geomaterial (soil or rock). The slurry inside the casing was bailed or pumped out, drilling continued to plan depth in the dry, a reinforcing cage was set in the partially cased borehole, concrete was placed and the casing was withdrawn. Some soil-slurry was undoubtedly trapped between the casing and the borehole wall, but the quality of that slurry was usually of no concern, because engineers of the period usually viewed drilled shafts as nothing more than deeply buried footings. Any unejected slurry would only serve to lower side shearing resistance, not yet considered in design, but would have no effect on end bearing. In the 1960 's the effectiveness of soil-slurry came into question because it was often ineffective in maintaining borehole stability at marginal drilling sites and because the economy of using shaft resistance, in addition to end

Jan 1, 1991

The objective of this study was two-fold: 1. To perform a world literature search pertinent to the analysis, selection and design of hoisting systems and their components and to codify the search into a literature compendium. The search resulted in a technical data file of over 1500 hard copy items. The compendium index was also suitable for establishing a comprehensive computer-loaded data base on the subject. 2. To develop a comprehensive system of guidelines to cover salient aspects of analysis, selection and design of electrical, mechanical and structural components of a hoisting system with a view to establish parameters for maximum safety and efficiency of operation, installation and revision of the system. 3. Considerations in the mine shaft layout, design, sinking, outfitting, testing and commissioning are set forth against the background of requirements of applicable parts of Code of Federal Regulations, Title 30.

Jan 1, 1988

By Jon C. Volkwein

The U.S. Bureau of Mines conducted a study to assess the impact of a directional water spray system on a unique single inlet dust scrubber on a low-profile mining machine. Tests were conducted in a full-scale model mine using a continuous mining machine. Tracer gas was introduced at the face and concentrations of the tracer gas were monitored at the face and in the return. Various parameters tested included mining position, depth of cut, brattice setback, and combinations of scrubber, and directional or symmetrical sprays. Results indicated that either water spray configuration improved the ventilation effectiveness of the scrubber by a factor of 2 to 3. Average face ventilation effectiveness (FVE) values for all testing of the scrubber and directional sprays were 0.95, versus 0.37 for a standard 20-ft blowing curtain without sprays. No FVE values were less than the standard 20-ft blowing results.

Jan 1, 1989

By E. A. Mihok

A pilot plant for treating mine drainage by neutralization with limestone was designed, fabricated, and operated by the Bureau of Mines. The process consists of (1) producing a very fine (minus 400-mesh) limestone slurry, (2) mixing the slurry with the mine water, (3) aerating the resulting mixture to remove carbon dioxide and precipitate iron, and (4) separating the solids from the liquid by sedimentation. A mine discharge of pH 2.8, containing 1,700 ppm total acidity, 36 ppm ferrous iron, and 360 ppm total iron, was treated at rates from 300 to 400 gpm. The treated water ultimately reached a pH of 7.4 and contained no detectable iron. The resulting sludge compacted to approximately one-third the volume of sludge from mine water neutralized with lime. The advantages of the limestone process, compared with the conventional lime process are as follows: Widespread distribution and lower cost of run-of-mine, indigenous limestone; simplicity of plant design; reduced hazards and plant operating and maintenance costs accruing from the use of a less reactive chemical agent; higher solids concentration of the precipitated sludge; and no deleterious effect caused by accidental overtreatment. While total process costs could not be determined, it is conservatively estimated that the cost of treating mine water by the limestone process is one-third to one-half the cost of the lime process, depending on the characteristics of the water to be treated and the availability and quality of limestone used.

Jan 1, 1968

By John A. Breslin

The Bureau of Mines measured respirable dust in coal mines by means of multiple dust samplers worn by persons moving about the mines. The measurements were made primarily to evaluate the effectiveness of certain dust-control techniques; however, for this report, the data have been analyzed to determine the precision of the personal dust-sampling measurements. A total of 68 sets of measurements were available from multiple personal samplers. The measured dust masses were less than 1.0 mg because the sampling times were short. The coefficient of variation calculated from these data was 18.5 pct, compared with coefficients of variation ranging from about 10 to 30 pct, which were calculated from the most precise and extensive data available for fixed-point dust measurements made in coal mines. From this comparison, it is concluded that dust samplers worn by personnel are as precise as fixed-point samplers in measuring respirabe coal dust; the wearing of the samplers by someone moving about a mine does not contribute significantly to the variation in measurements made with these instruments.

Jan 1, 1983

By B. J. Moore

This Bureau of Mines publication estimates that U.S. helium resources as of January 1977 were 710 billion cubic feet. These resources are broken down into four classifications, as follows: Helium in measured natural gas resources and in storage, 201 billion cubic feet; helium in indicated natural gas resources, 150 billion cubic feet; helium in hypothetical natural gas resources, 175 billion cubic feet; and helium in speculative natural gas resources, 184 billion cubic feet.2 Most of the helium in natural gas is presently dissipated as natural gas is consumed as fuel and for other purposes. Helium resources of the United States may also be divided into three other categories: (1) Depleting and undiscovered natural gas resources, 589 billion cubic feet; (2) nondepleting natural gas, 83 billion cubic feet; and (3) storage, 38 billion cubic feet, of which 37 billion cubic feet is federally owned and 1 billion cubic feet is privately owned. The United States does not have any significant leasehold rights in the depleting fuel gas resources. However, through the Department of the Interior, it does have leasehold rights to about 70 percent of the helium found in nondepleting natural gas, and it owns over 90 percent of the helium in storage.

Jan 1, 1979

By Louis Schaffer

The purposes of the research were to assess the applicability of new accident investigation methods to mining accidents, and to adapt those found to be applicable in the development of a new methodology for use in mining. The development included field testing, evaluation and modification of the new methodology. The report is in two volumes. Volume one discusses the background research, the design of a methodology for field evaluation, preliminary field tests in Utah mines, and work done to prepare for a formal, complete evaluation by federal government mine safety officials. This, volume 2, discusses the field evaluation done by the contractor and cooperating mines after the program was re-directed by the Bureau of Mines. It relates to research done during a period of approximately eighteen months, ending in April, 1983.

Jan 1, 1986

By Osvaldo A. Bascur, J. Pat Kennedy

"The profitability of mining and metallurgical operations continues to lag behind that of other industries despite the technical advances that have occurred in recent years. A primary reason for poor performing in this economic sector is the lack of integration of mining and metallurgical information. The mine cut-off grade can be optimized if mining and metallurgical engineers have the data to optimize milling for the type of ore in hand.Integration of mine transactions, stockpile inventories, process flows and assays is the key to determining the economics for each blast strategy in a mining site. Real-time information from manufacturing assets greatly affects the profitability and flexibility of a company, but gross errors from faulty sensors, inaccurate instrument calibration, or incorrect process models may corrupt actionable information. New software tools can model the process, sift through available data to identify and eliminate gross errors, and use statistical data reconciliation to create materially balanced and consistent readings. The reconciled data is unified for use by both monitoring and decision making processes.Creating a real-time information environment where upgrade and change are allowed is becoming an accepted methodology to accelerate and simplify the integration of applications such as production monitoring, equipment performance management and yield and inventory management."

Jan 1, 2003

By John M. L Lewis, Antoine F. Mulaba-Bafubiandi

In the irradiation of ore particles by microwave energy preferential absorption of this energy occurs in minerals having specific dielectric properties (Bradshaw, 1999). This leads to a selective heating of the various components of within the ore particle matrix. The result is a differential thermal expansion of such components and the resulting induced stresses become ideally concentrated at mineral grain boundaries (Kingman et al., 2000). Dependent upon the relative composition of absorbent, opaque and reflective components this may lead to enhancement of the inter-granular fracture of the ore particles in the initial comminution process (Fandrich et al., 1997). The potential is thus a greater proportion of free mineral grain within the fractured ore particle that may lead to improvement in subsequent mineral separation processes. As the composition of the ore dictates the result of such a treatment (Kingman et al., 2000) it is essential to evaluate the response of each ore sample on an individual basis. This paper reports on results obtained from two ore samples: the one being the zinc ore from the Black Mountain mine (South Africa) and the other a pre-concentrated sample of handpicked cobalt?copper mixed oxide ore from the Gecamines (DRC). The equipment and methods used to simulate the comminution of the samples and evaluate the level of the value liberation will be described along with the results obtained. Using King (1975), Kanellopoulous and Ball (1975), and Gungor and Atalay (1998), a comparison between the behaviour upon microwave irradiation of sulphide and mixed oxide ores leads to conclude that while the technique (i.e. microwave irradiation) is appropriate to the zinc sulphide ore, it does not seem to lead to exceptional results for the oxide.

Jan 1, 2005