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By R. L. Castro

The extraction of ore in panel caving operations also involves the coextraction of non-economic material, known as dilution. Forecasting, avoiding, and understanding the mechanics of dilution is critical to longterm planning. The models of dilution that are currently in use in panel caving operations are those by Laubscher (1994), who postulated that dilution is due to gravitational flow and depends on the height of interaction, the column height, and the draw uniformity. This paper presents a back-analysis of extraction and dilution behaviour at Codelco?s El Salvador and Andina Divisions, which use panel caving as a mining method. This analysis was conducted in order to review the existent dilution entry mechanisms and calculation methods. Three sources of dilution were identified. The first source is ?gravitational dilution?, which is due to gravitational flow and mixing as dilution migrates by gravity from a sector located above the panel under analysis. The second source of dilution is ?caving dilution?, which occurs when an air gap is formed in a large area and a sudden propagation of a hang-up occurs, inducing early ingress of waste into the production level from a level above. The third mode observed is ?lateral dilution?, which occurs when the panel is located next to an exhausted sector and the broken material enters the draw column early during draw. In the last two cases, dilution could migrate large distances horizontally. The mine data indicated that draw and caving strategy are key for long-term planning of large panel caves.

Jan 1, 2014

By P. A. Neill, O&apos

The caisson foundation for Vent Building No. 6, at the southern landfall of Boston's Third Harbor Tunnel, was installed inside an unbraced concrete cofferdam 250 feet in diameter and about 70 feet deep. As a spectacular engineering feat, the cofferdam itself certainly ranks first. However, the foundation work within was no routine caisson job and made its own contribution to the state of the art. Comprised of 5 ft. diameter caissons socketed 10 to 16 feet in sound rock, it posed serious challenges in logistics, space limitations and time, requiring development of unique, purpose-made equipment. The unexpected discovery of significant depths of shattered rock overlying acceptable sound rock, along with other surprises of lesser import, compounded the problems. How these challenges were planned for, met and overcome is the subject of this paper.

Jan 1, 1994

By J. Jerome MacDonald, Stephen B. Flewelling

"In the base metals industry of today, operations required to continuously improve productivity and reduce unit production costs. Consistent with these goals is the improvement of the muck haulage cycle from open stopes with the No. 12 Mine.Existing haulage from open stopes occurs with LHD units experience at Brunswick indicates that LHD units have a higher unit cost per tonne mucked than loader-truck combinations when the haulage exceeds approximately 150 m.The reintroduction of open stoping at Brunswick made existing loader-truck combinations unsuitable. It was recognized that application of recent continuous loading technology could potentially result in significant unit cost savings and productivity increases.A review of the technology resulted in the testing of the Voest Alpine AL60 loader in the open stopes located on the 1000m Le vel. Testing began in January 1988 and finished in late November 1988. The test will provide experience as well as quantified production and maintenenace parameters f or the purpose of quantifying production unit costs and system productivity, ultimately resulting in a decision to further utilize or not utilize the loader as a primarymucking unit at Brunswick .INTRODUCTIONBrunswick Mining and Smelting Corporation Ltd. produces approximately 10500 tonnes per day from the No. 12Mine located 35 km southwest of Bathurst, New Brunswick Mining is the largest (single mine) producer of zinc in the world having mined over 52 million tonnes since 1964. Many years of production still remain with proven and probable reserves in excess of 90 million tonnes. Reserves grade at an average of 3.65% Pb. 8.99070 Zn, 99g/mt Ag and 0.32% Cu . The operation employs 1446 people with approximately 800underground including operations and maintenance staff.In the No. 12 Mine two major mining methods are presently utilized . The primary method is Mechanized Cut and Fill (MCF) accounting for approximately 6500 tonnes per day from four major areas. The remaining 4000 tonnes per day is produced from open stopes with the majority of 2800 tonnes per day being produced by the 1000 level.The 1000Level contains 30 million tonnes segmented into open stope blocks 30 m high, 15m and 30 m on strike , with a transverse width of 10m to 50 m (Fig. I). Stopes are organized on a primary secondary stope arrangement along strike. Primary stopes are undercut and overcut full size with the back fully cablebolted. All stope and lateral development is driven 4.0 m high. All stopes to date are flat bottom (Fig. 2). Primary stopes are filled with cemented rock- fill , secondaries with conventional rockfill, Haulage from the stopes is primarily with 8-yard LHD units ; mucking is approximately 50% conventional and 50% remote. Haul distance to date has averaged approximately 150m with current stope haul distances of approximately 300m. Haulage is presently to a small jaw crusher situated on the 1000 Level. From August 1989on, haulage will be to a set of ore passes feeding to a large gyratory crusher situated on the 1200 Level. Start up of the gyratory crusher corresponds to an increase in tonnage from 1000 Level to 4100 tonnes per day."

Jan 1, 1989

Operations of the new copper smelter at Mount Isa commenced on the 6th February, 1953. Prior to this date a paper by R. V. Anderson (1953) outlined the smelter layout, proposed operation and expected performance.Nominally a 50-ton per day plant, output is currently at the rate of 85 long tons per day. In the first 3t years, of operation to 30th June, 1956, '71,339 tons of blister copper were produced. Operation has been stabilized' and a number of modifications to plant and operation have been made. This paper reviews the progress made to date.GENERAL LAYOUT AND EQUIPMENTThe general layout and equipment follows well-established practice as study of Figs. 1 to 4, and Appendix II will show. This diagrammatic representation of layout (Fig. 1) and flow sheet (Fig. 2) is intended for the purpose of brevity to take the place of a detailed plant description.Choice of site for the smelter was limited by the necessity for close proximity of the waste heat boilers to the powerhouse.Much excavation was necessary before foundation work commenced, but the plant is compact and further expansion is possible.CHARGE MAKE-UPRoaster charge initially comprised concentrates, purchased ores, revert materials, mine sulphide ore and barren sand. As purchased ore receipts increased and as domestic black rock oxidized ore became available, the latter two fluxes were eliminated from the charge. Fluxes at minus i-in. sizing are trucked from crushing plant or stockpile by motor lorries. Filtered mill concentrates are dropped from an overhead conveyor on to the storage patio floor adj acent to the two concentrate bins.

Jan 1, 1957

By J. H. Cloete, S. M. Bradshaw, D. K. Chibwe

"Modern tundishes have evolved as vessels to serve as the final step in refining of molten steel by removing inclusions and promoting thermochemical homogeneity.In this study the flow behaviour in a four-strand tundish was investigated by means of a ½-scale water model as well as numerical modelling. The numerical and physical models were used to characterize residence time distribution and calculate properties pertaining to tundish flow regime. Three different tundish configurations were investigated: a bare tundish with no flow control devices, a tundish with a turbulence inhibitor, and a tundish with both a turbulence inhibitor and a dam.The physical and numerical models showed that a tundish without flow control devices is prone to significant short-circuiting. A tundish with a turbulence inhibitor was shown to be successful in preventing shortcircuiting and provided surface-directed flow that might assist the removal of inclusions from the melt. However, it was also observed that the upward-directed flow caused the maximum turbulence kinetic energy near the surface to increase dramatically. The potential for slag entrainment should therefore be considered during the design and operation of tundishes with turbulence inhibitors."

Jan 1, 2015

The presence of undetonated explosives in a rock face after firing constitutes a hazard which can result in a fatal or serious accident if undetected before drilling commences.Most accidents caused by drilling into explosives are the result of failure to examine the working face properly and lack of care in subsequent drilling operations; it is not always realized that explosive may be left in the butts of holes or may be driven into cracks or crevices in the rock.Obviously the best means of preventing accidents of this nature is to adopt methods of drilling, charging and firing which will reduce the hazard to a minimum.There are three ways in which undetonated explosives may be left in the rock face, namely:1. Misfires.2. Cut-off shots.3. Incomplete detonation or failure of the charge to propagate throughout.MisfiresNo attempt will be made in this paper to deal fully with the question of the prevention of misfires, as apart from the fact that most of the preventive methods are well known, it is believed that if the miner receives his explosives and means of initiation in good order, and proper care is taken in their use, the failure of a charge to fire will be a rare occurrence. The provision of central fuse capping stations on mines has gone a long way towards the elimination of misfires caused by the mishandling of safety fuse anddetonators.

Jan 1, 1956

A pyrite flotation concentrate assaying 0·065 oz/ton gold was investigated to determine the mode of occurrence of the gold. Ftom cyanidation and acid leaching experiments it was concluded that O· 02-0· 03 oz/ton goldwas in solid solution in pyrite and the balance of the gold was finely dispersed as grains throughout the pyrite.Autoradiography and electron-probe examination of possible areas in briquetted concentrate did not detect any grains of gold. Neutron irradiation and gamma spectrometry confirmed on 100 mg samples the fire assays obtained on 1 assay ton samples.Oxidation to hematite and reduction to magnetite at 600°C made practically the whole of the gold extractable by cyanide. This technique could be of major significance for gold ores which contain a substantial amount of "refractory" gold.

Jan 1, 1969

IN this paper it is proposed to present certain features of operating practice at the Port Pirie works of The Broken Hill Associated Smelters Proprietary Limited, in the smelting in a blast furnace of lead ores carrying a high concentration of zinc.The subject matter selected for presentation is:(i) The effect of zinc on lead blast-furnace operations, together with an outline of the methods adopted in actual practice in dealing with the zinc problem.(2) Development of the height of tuyere above the bottom of the jacket, and the extension of the tapping interval and the combined effect of both factors upon the lead content of blast furnace slag.It will later be seen that one phase of the investigation carried out in connection with the effect of zinc on blast-furnace operations was largely responsible for the decision to increase to a marked degree the height of the tuyere above the bottom of the jacket.Due to this interdependence, it has been considered advisable to deal with the two features above mentioned.The high concentration of zinc in the lead-bearing materials smelted at Port Pirie has brought with it those troubles which are associated with zinc in the smelting operation.In the early stages of the history of The Broken Hill Associated Smelters Pty. Ltd. the trend towards increasing cost for labour, fuel, and stores, etc., rendered it necessary to increase the lead content of the charge, and reduce to a minimum the quantity of barren flux used.

Jan 1, 1931

By Frank A. Peters

This evaluation of the sulfurous acid process for the production of metallurgical-grade alumina is the first of a series concerning various known methods of extracting alumina from low-grade aluminous materials and was undertaken to permit comparison with other processes. In the process with which this Bureau of Mines report is concerned, calcined clay is leached under pressure with 20 percent sulfurous acid and then filtered to remove a siliceous residue. The aluminum sulfite remaining in the filtrate is precipitated as a monobasic sulfite which, together with iron sulfite, is then decomposed under pressure to form a crude alumina hydrate and iron oxide. This intermediate product is purified in a modified Bayer process to yield alumina suitable for electrolytic reduction to aluminum. A study estimate (16)3 is used for these evaluations. The capital cost, not including land, is estimated at $95,701,600 on a Marshall and Stevens index basis of 240.0 for a plant producing 1,000 tons a day of alumina or 350,000 tons a year. The annual operating cost for this size plant is estimated at $22,831,100 or $65.23 per ton of alumina using clay at $1 per ton. Depreciation and local taxes are included in the estimate, but profit and Federal taxes are not considered.

Jan 1, 1962

By C. B. Manula

This report is volume 9 of an eleven-volume final report for a Grant completed by The Pennsylvania State University for the Bureau of Mines to develop a dynamic general purpose computer simulation model for underground coal mining. The simulator developed under the Grant has a wide range of possible applications in the design and evaluation of underground coal mining systems from the standpoints of health and safety, productivity, and mine environment. The Master Design Simulator consists of nine free-standing function-oriented subsystem models linked together into an integrated, comprehensive model of an entire mining system. The subsystems comprising the Master Design Simulator include geology and reserves, methane generation, roof support, subsidence, production, rail haulage, cost, water generation,' and ventilation. This volume of the final report describes the Water Generator subsystem of the Master Design Simulator. This subsystem simulates the flow of water through a typical watershed-reservoir combination to predict mine drainage quantities in underground workings. The model is a integrated one which handles both the unsaturated and the saturated flow domains as a unified system. The solution of the model is achieved by numerical tech¬niques of finite difference method coupled with finite elements analysis. The Water Generator can be operated either as a free standing model or in connection with the Geology and Reserves subsystem as a part of the Master Design Simulator. A case study is included illustrating the application of the model to actual data. The appendices of the report constitute a user's manual for the program, and include a program listing.

Jan 1, 1975

By J. A. Finch, Zhenghe Xu, Z. A. Zhou

Whether the microscopic bubbles generated by the high energy dissipation and high feed-slurry velocities in new flotation cells are stable in solution and what roles they play in particle collection are of great interest for the flotation of fines. Experimental work has shown that, with fast flow of liquid through an orifice, stable microscopic bubbles can form in flotation systems by hydrodynamic cavitation in the absence of gas supersaturation. The minimum flow rate at which bubble formation is initiated varies from 8 to 15 m/s, depending on the design of the cavitation tube, which is within the range of feed slurry velocities of some newly developed cells. Feed aeration enhances the formation of micro-bubbles by hydrodynamic cavitation. The study suggests that the fast flotation kinetics of some new cells may be attributed in some degree to the formation of microscopic bubbles

Jun 21, 1905

A reliable laboratory method to characterize the response of free gold particles to flash flotation conditions has been developed. The test has been performed on free milling gold ores as well as synthetic ores, using either a gravity concentrate or gold powder as the gold source, to assess the floatability of gold particles. Trends in free gold flotation kinetics, as well as size and milling effects, were identified for gold recovery based on the different feed types, reagent dosages, and residence times. It was shown that the ultimate recoveries and kinetic trends of gold particles from the gravity concentrate could be enhanced with increased dosage of collector, potassium amyl xanthate. Interestingly, in comparison to gravity-recoverable gold, recovery from pure Au powders was better in collectorless flotation, and cumulative recovery decreased with higher levels of collector addition. Improved coarse particle recovery appeared linked to increased collector additions for both the gravity concentrate and the pure gold powders. In general, coarse gold particles demonstrated slower kinetic rates thaen the fine or intermediate components incomparable tests.

Jan 1, 2015

By H. P. Kritzinger, T. C. Kingsle

"Rotary kilns are used for a variety of mineral processing operations. Hatch makes use of a kiln model developed from first principles to evaluate designs for its clients. This tool has been applied to a variety of pyrometallurgical applications, including ferrovanadium, nickel carbonate, nickel laterite, iron ore reduction, and spodumene (lithium) production.This paper illustrates the application of numerical optimization techniques in combination with the kiln model in the interrogation of a generic iron ore reduction process. The fundamental modelling concepts are explained, followed by a description of the optimization approach. The results show how the combination of the two methods, with modern computing power, can generate a large number of viable design and operating candidates."

Jan 1, 2015

By R Soong

The Wharekirauponga (WKP) epithermal gold-silver deposit is located about 10 km north of Waihi in the Hauraki Goldfield east of Auckland. WKP is notable for its sheeted to stockwork quartz veins hosted in rhyolitic rocks, that have attracted exploration interest as a low grade, bulk tonnage exploration target. Exploration between 1978 and 1993 included 23 diamond drill holes totalling 5505 m. The mineralisation occurs in a rhyolite dome complex (Miocene), consisting of flow-banded rhyolite and rhyolitic pyroclastic rocks, with minor andesite. Fifty four drill core samples were collected from six exploration drill holes and analysed by semiquantitative XRD techniques. Hydrothermal alteration of the rhyolitic and andesitic rocks has developed secondary mineral assemblages of quartz, adularia, albite, chlorite, illite, sericite, interlayered illite-smectite, carbonate, halloysite and pyrite. The main zone of mineralisation is characterised by quartzadularia alteration, which grades oiItward to quartz-clay (illite, local halloysite and/or interlayered illite-smectite) alteration. Propylitic alteration is present at the southern end of the drilled area of the prospect. Depth-related changes in mineralogy include a decrease in illite and appearance and increase in quantity of chlorite with increasing depth, and the localisation of halloysite and interlayered illite-smectite in the shallow parts of some drill holes.

Jan 1, 2001

By Drupad B. Desai, Bruce Chang, Jay L. Merritt

This paper reviews difficult ground conditions encountered along the alignment of large-diameter sewage tunnels for the City of Los Angeles. It offers a detailed one-of-a-kind solution to a unique problem for shallow urban underground tunnels in a seismic zone, using state-of-the-art materials, technology, and equipment. The following topics are discussed in detail: • Geotechnical Design consideration for tunnel traversing through an active fault zone. • Seismologist evaluation of possible fault-offset problems where a fault zone crosses the alignment, in the event of an earthquake of moment-magnitude 6.5. • Probability analysis regarding "recurrence interval." • Structural design of tunnel subjected to displacement, subjected to axial elongation or shortening, and capable of rotational freedom. • Special joint details.

Jan 1, 1989

By E Villaescusa, A Kusui

The depth of underground mines has been increasing due to several factors, such as material demand growth and improving technology. It is a challenge to understand severe failure mechanisms such as spalling, which is caused by induced high stress. To investigate this failure mechanism under high stress conditions, intact rock blocks with small size tunnel were tested at the Western Australian School of Mines (WASM). The range of rocks, such as sandstone and granite, were drilled (200 mm hole) and tested using a servo-controlled 450 t capacity INSTRON machine. The acoustic emission sensors were attached to the block mass during the experiment to investigate the seismic response prior to spalling failure. The effect of the ground support system towards the failure was also examined by comparing the seismic response of the various ground support schemes.CITATION:Kusui, A and Villaescusa, E, 2016. Seismic response prior to spalling failure in highly stressed underground tunnels, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 689–694 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By J. Kortnik

Mining has always been, and still is, an activity which is vital for the development of mankind. This is because a predominant proportion of industrial products and energy is produced from raw materials that are excavated using mining methods. The need for environmental protection and the close proximity of urban settlements and mines in Slovenia have necessitated ever stricter requirements for the excavation of raw materials and introduction of more environmentally-friendly mining methods. This is achieved by implementing a closed eco-technological cycle of raw material excavation within the scope of mining methods. In surface mines, this means prompt reclamation of the degraded surfaces, and in underground mines the use of methods with minimized impact on the surface and on underground water. In underground mines, mining methods with backfilling installed according to the multi-barrier disposal system are the only ones that allow the possibility of implementing a closed eco-technological cycle of excavating raw materials. Various industrial waste materials are also included in the backfilling material in order to improve its strength properties. Backfilling consisting of several different waste materials is produced in the form of a composite. Before backfilling consisting of waste material composites can even begin to be used in mines, an environmental impact assessment of its suitability for mines needs to be made. In order to make such an assessment, a pressure leaching cell and an open diffusion cell were made, i.e. a pressure leachate test and a sorption test were developed. The two cells and the developed tests will enable the study of the influence of backfilling on the environment if installed in an underground mine. The pressure leachate test performed with the use of the pressure leaching cell simulates leaching from a consolidated backfill during the seeping of underground water into the backfilling material under pressure. The sorption test performed in an open diffusion cell simulates the flow of eluate leached from the backfilling material through the rock/geotechnical barrier. The two tests were developed in such a manner that they are performed either in succession immediately after one another, but they can also be performed separately at a certain time interval. The paper presents an assessment of the suitability of backfilling made of different waste material composites. Keywords: environmental impact assessment, multi-barrier disposal system, waste material composites, pressure leachate test, sorption test, pressure leaching cell, open diffusion cell, eco-technological cycle, mining methods with backfilling.

Jan 1, 2003

By Fred E. Fiedler

A demonstration project investigated the feasibility and effectiveness of two management interventions, Management Training and Organization Development, for improving productivity and safety in metal/non-metal mines. The management training intervention was conducted at Texasgulf Company's trona mine, using three nearby trona mines as controls. The Organization Development intervention took place in Hecla Mining Company's Lucky Friday mine, using the highly similar Star Mine as control. The two management interventions seemed about equally effective in reducing accidents and injuries, as well as MSHA citations, and the management intervention at the Texasgulf mine also appeared to have led to substantial improvement in mine productivity.

Jan 1, 1983

By A. de Meyer

Zincor experienced significant losses of metallic zinc to dross during the melting of zinc cathode plates. In this study five factors affecting the formation of dross are identified from literature. Dross from both the plant and that produced in the laboratory is characterized using XRD. An empirical model, based on laboratory-scale investigations, was developed to quantify the effect these factors on their own or in combination with each other has on the amount of zinc loss to dross. The predictive capability of the model is evaluated based on the laboratory-scale investigations. Predictions are made on the effects that heat treating the cathode plates prior to melting, increasing the cathode plate thickness, washing and drying prior to melting, and loading the furnace in a controlled manner have on the amount of zinc loss to dross. Keywords: Zinc, dross, optimize, smelter, yield, losses, empirical, model, XRD.

Jan 1, 2010

By P. M. Daling

This report, sponsored by the U.S. Bureau of mines, presents the results of an examination of available safety assessment methods and their applicability to the mining industry. Standard safety analysis procedures were evaluated for relevance to the mining industry and mire operators' needs. The major safety concerns and needs of mine operators were identified through telephone contacts and a visit to a mine site. Safety assessment methods recommended for use by the mining Industry were also investigated for potential adaptability to a user-interactive computer system. Several methods were determined to be Immediately transferrable to the mining industry; these techniques are described in a user's handbook produced as part of this project. The handbook gives detailed instructions on appropriate application and utilization of these methods. Other techniques will require further development before they can be effectively transferred to the mining Industry. These methods, their advantages, disadvantages and resource requirements are described in the final project report. Recommended system safety analysis techniques facilitate an evaluation of total mine system safety and a determination of safety-related interactions among systems within a mine. The application of these methods to the mining industry could result In reductions in both the frequency and the consequences of mine accidents.

Jan 1, 1983