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By J. S. Phipps, D. R. Skuse

"Aqueous slurries of fine particle size calcium carbonate are used in a diverse range of end-use products. In many cases, the end use performance is governed by the fineness of the particle size distribution. In turn, the maximum fineness of the particle size distribution is primarily determined by the efficiency of the dispersants used during the production process. Here we report the sodium polyacrylate mediated dispersion of calcite and kaolin. Firstly, we report a colloid chemical description of the dispersant behaviour and attribute the stabilisation to both steric and electrostatic mechanisms. Secondly, we describe how the understanding developed in the first part of the study can be utilised to produce more effective dispersant systems which allow the preparation of, added value, higher solids and finer particle size calcium carbonate slurry products.INTRODUCTIONConcentrated aqueous dispersions of fine, particulate kaolin and calcium carbonate are used extensively in the coating of paper. Currently more than 10 million tonnes of such dispersions are used annually in North America and Europe alone, and worldwide consumption is growing at around 3% per year.Generally, paper is coated to enhance its visual appearance. Firstly, coatings improve the optical properties of the blank sheet, i.e. its brightness (or whiteness), opacity and gloss. Secondly, they improve the physical properties of the sheet that will influence the quality of the final printed image, such as surface smoothness, porosity, and ink absorption. In order to achieve these aims economically, paper coatings must be applied at high speeds and with minimal drying costs.Kaolin and calcium carbonate pigments are the major components of most paper coating formulations. Particle size distributions are chosen to maximise the optical benefits that they confer, and to allow the production of formulations that can be easily pumped, handled and coated. Coating pigments are generally produced and transported as slurries, and so it is important to obtain dispersions with the highest possible solids content, both to minimise drying costs and the quantity of water transported. This paper examines the dispersant technology that is currently used in the industry and the mechanisms by which it functions, and illustrates ways in which this technology has developed and could be further improved in the future."

Jan 1, 2003

By T. A. Kirk

"The Linatex Dewatering System is a compact, low capital and maintenance cost matched set of processing equipment suitable for solids recovery, dewatering and desliming sand-like materials in the range 3/8"" to 325 mesh ( 1). Since its introduction to North America in the early 1970s, various elements of the system have been successfully installed in a variety of applic11tions both in Canada and the U.S.A. Examples are chosen from the coal, fine aggregates, iron ore and industrial minerals industries.II. Elements of the Linatex Solids Recovery System(A) The Linatex Separator (See Fig.I)The Linatex Separator, which is the heart of the system, is a specialized type of hydrocyclone designed to give a self regulating rope discharge product. It has the familiar cyclone geometry of cylindrical and conical section, tangential inlet, apex and vortex finder. The apex is enclosed in a unique patented automatic underflow regulator consisting of a collapsible rubber sleeve, and the overflow line is a siphon-leg, with a controlled air bleed valve. The siphon leg is adjusted so as to balance the internal negative pressure acting as a closing force on the underflow regulator against the mass of slurry pressing down on it from the cone section of the Separator. With correct adjustment the underflow regulator will close drip-tight on a water only feed. As solids accumulate in the apex of the cone the weight of solids counterbalances the vacuum and cat1ses the underflow regulator to open and discharge classified material at a controlled density."

Jan 1, 1978

By F. G. Pickard

"The Strathcona Concentrator is one of three concentrators presently operated in the Sudbury District by the Nickel Division of Falconbridge Nickel Mines Limited. Production commenced in 1968 at a milling rate of 6,300 tons per operating day. Milling capacity was increased in 1970 to 7,200 tons per operating day. 7he ore contains iron, nickel and copper primarily in the form of pyrrhotite, pentlandite and chalcopyrite, while the common gangue minerals are feldspar, quartz, mica, talc and chlorite. The primary function of the concentrators is not only to liberate sulphides from silicate gangue minerals but also to separate pyrrhotite from the more valuable minerals pentlandite and chalcopyrite.SummaryExcessive wear on steel liners and blinding of grates in a regrind pebble mill of the Hardy Concentrator was corrected by the use of rubber liners and grates. The success of these resulted in their installation as original linings In all grinding units except the rod mills at the Strathcona Concentrator. The operation of rubber liners and grates when sized mine ore was used as the grinding medium was satisfactory confirming the Hardy Concentrator experience. Problems developed when the grinding medium was changed from sized mine ore to steel balls in that blinding of the grates occurred. This resulted in conversion of the units to overflow type mills through removal of the grates. The lifter bar life was significantly increased by using higher lifter bars.Grinding PracticeOperation of a regrind pebble mill commenced in 1961 in the Hardy Concentrator. This operation demonstrated the economic feasibility of using sized mine ore as grinding medium. Problems caused by excessive wear of metal liners and grates combined with blinding of the grates were corrected in 1964 through installation of rubber liners and grates."

Jan 1, 1974

By M. A. T. Cocquerel, M. G. Burcher, J. G. Eacott

"This paper is based on work carried out to examine the feasibility of increasing the smelting rate of two oil-fired calcine-charged reverberatory furnaces . Various constraints prevented physical enlargement of the furnaces and so allention turned to the use of preheated combustion air or oxygen enrichment to bring about the required increase in production. As a basis for further calculation, a furnace heat balance is determined using the smelting metallurgy and combustion data for the current operation. Various other operational data are tabulated and compared with similar data from other operating copper smelters.New heat balances are derived for various levels of combustion air preheat at the current smelting rate, and the effects of air preheat on oil consumption, combustion air and waste gas volumes are tabulated. These data are used to demonstrate the level of increased throughput which would be obtained before constraints in existing gas handling equipment are encountered.A similar exercise for oxygen enrichment is set out, with oil consumption, combustion air and waste gas volumes at various enrichment levels tabulated for the current smelting rate. These data are in turn used to demon strate the smelting rate which would be achieved at a practical level of enrichment before waste heat boiler and gas handling constraints are encountered.The energy implications for both cases are summarized."

Jan 1, 1980

By G. Castillo, G. Thies, C. Bissue

"Newmont’s Phoenix mine began gold production in 2006 and has since experienced difficulties in reaching design throughput due to variability in ore hardness. In 2010, a debottlenecking study was conducted, generating multiple options to increase throughput of the existing Phoenix Mill through a series of phased expansions. An option was selected in early 2011 to increase the current concentrator capacity by approximately 25% through two phases. The first phase of the project was the installation of a secondary crushing and screening circuit in order to reduce SAG feed size from a P80 of 6 inches to 2 inches. The second phase focuses on downstream unit operations in the concentrator aiming to relieve existing bottlenecks such as flotation. This paper analyzes the design, commissioning, and operation of a recently installed SAG Pre-crush circuit (Phase 1). The design discussion will focus on concept development, flowsheet design, design criteria, process models, design limitations, and control philosophy. Discussion on the integration of operating data and flowsheet into design during engineering will be presented. Actual operating data of major equipment is gathered from the commissioning and operation of the circuit and compared to the initial design models. Conclusions will be drawn on equipment design factors used during engineering, and the success of model integration into operation. The paper concludes on the successes of the project and where improvements can be made during design.INTRODUCTIONThe Phoenix mine is located within 25 miles of the town of Battle Mountain, Nevada. The mill processes copper, silver, and gold ore. Production started in 2006 and has since struggled to reach the designed tonnage of 33,600 stpd due to the variability in ore hardness. In production year 2010-2011, the Phoenix mill average daily throughput of 33,000 stpd. In 2008, a larger primary crusher (FLS 60/89 gyratory crusher) was installed to operate in parallel with the originally supplied primary crusher (Metso 50/65 gyratory crusher), in order to provide additional capacity upfront in the circuit so that the primary crushing circuit could keep up with the mill.In 2010, a mill debottlenecking study was completed to assess the various options that could be implemented to increase the current mill capacity. The study was completed in late 2010 and contained four tonnages that Phoenix could achieve through various expansion scenarios. A single option was selected in early 2011 to increase the mill capacity from 33,000 stpd to 48,000 stpd. The selected option looked at a four phased expansion that would increase the capacity of the Phoenix mill and address the mill’s current bottlenecks:"

Jan 1, 2013

By Camilo I. Morales

At present, many open pit mines are designed to operate and feed a processing plant with ore characteristics that were defined in the design of the mining business, which are not necessarily the same as those found during the operation. This is the case of Spence Mine, whose design was conceived considering the extraction of primary iron sulphides, however, with the extraction of such minerals, it was necessary to mine and process minerals with secondary enrichment. Then it is necessary to work with a greater selectivity degree in the operation, however the capacities of the mine and operational resources were not designed for this. Thus, the changes in the general conditions of the operation should be considered as a series of geometallurgical restrictions to achieve production schedules. The above facts impacts directly in the short-term planning, whose task is to achieve the goals proposed in productive long-term planning because the company does not have the tools necessary to plan with the degree of selectivity required, so it should be done by hand, discretizating and scheduling the production polygons. Naturally, this manual process in the short term planning deviate sharply from the areas in operation, opening new fronts and benches that result in a significant increase in costs and a loss of the value of the mineral resource to miss the exploitation sequence. The research presented in this paper is related to the development of a mathematical programming model to support the short-term planning, so as to integrate the challenge of selective mining with geometallurgy restrictions, to provide the planner a guide to construction of the polygon of exploitation, guiding the fines production to achieve the goals of the proposed production plan.

Oct 1, 2009

By R. E. Jones

"INTRODUCTIONCominco Limited, one of the free world's largest producers of lead and zinc having majority ownership in Pine Point Mines Ltd., operates a mining and concentrating complex on the south shore of Great Slave Lake in the Northwest Territories. The mine was developed in 1963-64 and commenced operation late in 1964 with the direct shipping of high grade ore. The concentrator started production in November 1965 at a design throughput rate of 5,000 T.P.D.In 1968, the Company acquired the mining rights to a large ore body located to the east of the Pine Point claims and subsequently an expansion program at the concentrator raised the throughput rate to 8,000 T.P.D. The expansion included a third grinding circuit and additional flotation capacity. The flotation capacity took the form of a completely separate circuit from the existing flotation plant. In 1972, the zinc flotation circuits were modified to include two additional stages of cleaning. This change concluded the ""Pre-Consolidation Flotation"" circuit at Pine Point.In considering these modernization programs of the flotation section of the concentrator, the philosophy was to utilize the largest available machine having proven and modern technological changes. With the advent of the large capacity machines available in the late 1970's, it was considered a suitable time to revamp the entire flotation circuit at Pine Point in 1979. The Circuit Consolidation Project was developed to meet this requirement."

Jan 1, 1982

By Christian Ottergren

"Modern crushers offer double the capacity and performance of classic models. In some cases crushers are replacing SAG circuits, the advantages being smaller investment and lower operating costs. Is Tertiary crushing replacing SAG milling as the most optimal grinding method? Does tertiary crushing achieve the same outcomes as SAG milling? This paper considers these questions and provides an assessment of the strengths of tertiary crushing in comparison to other methods. Two case studies: a Scandinavian Platinum Project, and CODELCO, are also presented.INTRODUCTIONResearch on basic parameters in crushing started at Chalmers University in 1993. The initial project has developed and now involves several related areas. Modelling in accordance with new theories has resulted in modified designs which have been implemented and shown improved performance at operating mines in Chile. Sandvik Rock Processing is a partner and a major supporter of this program. This paper will cover some of the achievements and give references to the research reports.The mining industry has for many years mainly used crushers of the traditional 7’crusher concept, with thread adjustment of the setting and with a spring or hydraulic overload system. This machine type has been the workhorse for crushing throughout the world. The explanation for the dominance of this machine is that it has a reliable mechanical design. It can be estimated that about 2000 crushers of this design are still at work.Technical development in all fields utilizes new technologies like high-strength materials, advanced control systems and more extensive use of hydraulics. This is the reason why today’s motorcars, for example, are more efficient, more reliable and more environmentally friendly than their predecessors in the 1950s. Crushers are no exception. This paper will cover experience from the latest-generation cone crushers developed by Sandvik Rock Processing.The parts in this paper related to the CODELCO ElTeniente copper mine in Chile are coauthored with responsible personnel from the operation at the mine."

Jan 1, 2004

By Jiwen Zhang

Abstract Due to the blocking nature at the crusher station, straightforward application of queuing theory cannot give realistic results on the performance of a shovel-truck-crusher system. A new queuing network-based model has been developed in which, the crusher station is formulated into an open queuing network and the whole shovel-truck-crusher system into a closed queuing network with the crusher station as a special server. To fully account for the influence of blocking, Monte Carlo simulation is first used to obtain the performance parameters of the open queuing network for the crusher station. The closed queuing network for the entire system is solved by applying the Extended Summation Method (ESUM), in which the crusher server is described by the simulation results. The model has been applied to the Yuanbaoshan open-pit coal mine to analyze its shovel-truck-crusher system and to improve its efficiency. Results indicate that the blocking at the crusher is an important factor affecting the crusher output and in turn the efficiency of the whole system, and prove that the model can realistically describe the behavior of such systems.

Oct 1, 2009

By Jiwen Zhang

Due to the blocking nature at the crusher station, straightforward application of queuing theory cannot give realistic results on the performance of a shovel-truck-crusher system. A new queuing network-based model has been developed in which, the crusher station is formulated into an open queuing network and the whole shovel-truck-crusher system into a closed queuing network with the crusher station as a special server. To fully account for the influence of blocking, Monte Carlo simulation is first used to obtain the performance parameters of the open queuing network for the crusher station. The closed queuing network for the entire system is solved by applying the Extended Summation Method (ESUM), in which the crusher server is described by the simulation results. The model has been applied to the Yuanbaoshan open-pit coal mine to analyze its shovel-truck-crusher system and to improve its efficiency. Results indicate that the blocking at the crusher is an important factor affecting the crusher output and in turn the efficiency of the whole system, and prove that the model can realistically describe the behavior of such systems.

Oct 1, 2009

By D. Turk, P. Lefebvre, J. Sargent, K. Major

"Stillwater Mining Company recovers platinum and palladium from their Nye Montana operations. Mining first commenced at Nye in 1987 at a milling rate of 455 t/d (500 stpd). The plant was expanded in 1995 with the replacement of the ball mill and the addition of flotation capacity to increase mill throughput to 1820 t/d (2000 stpd). The applications for platinum and palladium have increased in recent years particularly in the automotive industry and the supply of the metals has remained fairly constant with prices for both platinum and palladium increasing and remaining very stable. In 1997 Stillwater initiated the expansion of their operations to increase production of platinum and palladium. Projects were initiated to expand production at the Nye operations, increase smelter and refining capacities at the Columbus facilities and to develop the extension of the orebody at the East Boulder site. This paper will discuss the changes at the Nye concentrator which were completed to increase production from 1820 t/d (2000 stpd) to 2730 t/d (3000 stpd) and eliminate some of the process bottlenecks.IntroductionStillwater Mining Company is mining the only significant primary source of PGMs outside of South Africa. The J-M Reef is an extensive mineralized zone containing PGMs that has been traced over a strike length of about 43 km (27 miles) in the southwest region of Montana. Stillwater controls the entire PGM mineralized zone.Johns-Manville geologists discovered palladium and platinum in the early 1970’s. In 1979, a Manville subsidiary entered into a partnership with Chevron, USA to develop the project. In 1 987, operations at the Stillwater property near Nye, Montana commenced."

Jan 1, 2000

By V. A. Cooney

Introduction For the compliment paid me by the Institute's kind invitation to speak this afternoon, I am very grateful. For the astuteness of your officers in determining that encouragement should be given, through the medium of a paper, towards accident prevention in coal mines, the coal industry cannot but be grateful. Some day, and let us hope soon, a more developed paper may supersede this one. In the meantime, if some knowledge of what one coal field has been able to accomplish, even in the initial stages of its effort, will induce other fields and other sections of the mining industry to launch similar experiments, much good will result and a more developed dissertation on the subject will make its appearance.

Jan 1, 1946

By Sibabrata Patnayak

A major problem during shovel excavation in oil sand mining is the variability of material diggability resulting in varying mechanical energy input, cycle time, energy consumption, wear, and stress loading of shovel dipper-andtooth assembly. This variability greatly influences the mining equipment performance. Different approaches such as empirical ground classification systems, analytical methods, and performance monitoring that are used to determine diggability of groundmass are reviewed. It is concluded that the existing empirical diggability classification systems are site specific and mostly based on rock and therefore, cannot be used for shovel excavation in oil sands. The O&K wedge test is a simple and useful index test for establishing correlations between cutting resistance and geotechnical properties of coal and other soft rock overburden material. But this test, when performed on small size unconfined oil sand samples, may not give reliable indication of diggability because of the boundary conditions on the specimens. Digging predictions made using analytical and laboratory digging models need to be verified with performance monitoring data to better understand the digging effort in oil sands.

Apr 1, 2005

Taking Up the AFN Corporate Challenge ?? About MAC ?? ?Towards Sustainable Mining? ??Mining and Aboriginal Peoples Framework ?? AFN-MAC Relationship ?? A New MAC Northern Initiative ??Social and Economic Progress ?? A View to the Future?

May 1, 2009

By Patrick M. Y. Waye

As mines utilise more automation in mining operations to improve efficiency and safety, a corresponding increasing demand is placed on the transport of information. Some of the recent technological advances in underground telecommunications are described for various data, voice and video applications. In particular, two new innovative underground communication systems are described, one with high speed data at 30 Mbps and the other for mine wide evacuation and safety applications. The high speed data system incorporates stateof- the-art data networking technologies and the existing leaky cable narrow band radio channels. The new system provides over the same basic infrastructure the high speed data network at 30 Mbps TCP/IP Ethernet with 100 Base-T interconnection, plus 32 narrow band radio channels. The second system is a system for mine wide evacuation with "through the earth" communication infrastructure. Emergency situation can be communicated to and from all the miners within seconds through a central control location. The technology involved does not require leaky cable or any other similar transmission media installation. Many applications are possible, including warning miners of emergency situations, mine rescue operation to communicate with trapped miners, regular reporting from miners working alone.

May 1, 2001

By A. J. Magerowski

Energy consumption is a main component of operational costs for the comminution of mineral bearing ores. Reduction of energy consumption has continually been addressed by mill operators and equipment manufacturers. Our testing over the last few years strongly indicates that the overall energy needed to comminute a mineral bearing ore can be brought down significantly, if more work is done in the crushing stages of the comminution circuit.This calls for uprating of the conventional crushing equipment into a new generation of crushing equipment to produce finer products. This will result in the emergence of new ideas in the design and application of equipment and corrnninution circuits.

Jan 1, 1986

1. Scope of project In South Australia, the Western Mining Corporation operates the underground copper and uranium mine Olympic Dam. The deposits concerned are enormous (approx. 570 million tons) and have been mined since 1988. At the end of 1996, Western Mining decided to increase the capacity of Olympic Dam from 85,000 t.p.a. to 200,000 t.p.a. Planning At this point, SIEMAG-GHH Mining Division was asked to carry out a study including the planning of the entire shaft hoisting operations and calculation of the investment and operating costs involved. With the assistance of Australian partners, the study was completed within a single month. The principal points of the study were as follows: ? Planning of the hoisting machinery for 1,375 tons per hour from the surge bin below ground to the discharge bunker above ground. ? Optimisation of the hoisting operation with regard to maximum availability and safety. ? Optimisation of the time schedule to minimise the financing costs of the investment. ? Statement of the investment and operating costs. ? Comparison of the technical data, design and availability of the planned machinery with reference plants. Construction of the plant Following acceptance of the study, the decision was taken within a few weeks to implement it. SIEMAG Transplan GmbH received the order to execute the entire shaft hoisting machinery on the basis of the study submitted. The decisive factors in placing the order were: ? Convincing study, adherence to schedule ? Reference to comparable plants elsewhere ? Reference with regard to availability.

May 1, 2003

By K. Duckworth

"Comparative tests of the conventional Turam survey configuration and an alternative configuration called the Tx-parallel configuration were conducted at the McClean uranium deposit in northern Saskatchewan. Both configurations detected the known conductors at the site but the alternative configuration demonstrated its particular effectiveness in the reconnaissance role. This effectiveness results from the fact that the length of Tx-parallel traversesis controlled only by the length ofthe transmitter and not by its power. In addition, the field strength along a Tx-parallel traverse is uniform unless the traverse approaches the end of the transmitter. This freedom from a strong gradient of the primary field along a Tx-parallel traverse eliminates the gradient-induced anomaly shifts which are common in conventional Turam operations. The locations of each of the McClean conductors, that were obtained from the Tx-parallel profiles, werein better agreement with the locations known from drilling than were the locations obtained from the conventional profiles. However, the conventional profiles gave better estimates of the conductance of the targets.The results obtained over the McClean conductors indicate that current gathering has a strong influence on Tx-parallel results with consequent unpredictable influence on depth and conductance estimates. However, the depth estimates obtained from the Tx-parallel data appear to be in good agreement with the known depth of the McClean south conductor. The Txparallel results provide shallow estimates of the depth of the McClean north conductor which suggest that the fracture zone above the graphitic conductor is a significant conductor as was also suggested by earlier VLF surveys.Comparison with the wide range of published geophysical test surveys for the McClean site showed that apart from the work required to lay the transmitter, the Tx-parallel mode of operation of a Turam exploration system is operationally very similar to a moving source Slingram type of survey. At the McClean test site, it provided the same depth of exploration as a conventionally deployed fixed loop transmitter."

Jan 1, 1988

By Anand Raman, Steve Dixon

"Industry practice for defining acid consumption is based on column tests. Freeport-McMoRan (FCX) has adopted a practice of leach operation using a defined pH or gpl acid in the raffinate based on the column test program. The acid consumption can be easily calculated using the leach operating parameters of SX-EW efficiency, lift height, bulk density, irrigation rate, primary leach cycle time, acid cure dose and acid concentration in the raffinate. The column leach test results and a column to heap efficiency factor are used to define the copper extraction. The maximum acid consumption will occur when zero acid is contained in the raffinate. The term Net Extracted Value (NEV) is to define the maximum extracted value relative to acid consumed.This paper presents the operating logic and method used to calculate acid consumption based on column tests and operating parameters.INTRODUCTIONIndustry practice for defining acid consumption is based on column tests. This may result in lower economic value if the tests are not designed to account for operating variables. A test program is described to assist in defining lift height and the influence of gangue on solution chemistry. A simple spreadsheet is presented that will use the leach operating parameters of SXEW efficiency, lift height, bulk density, irrigation rate, primary leach cycle time, acid cure dose and acid concentration in the raffinate to calculate the acid consumption for a proposed operation.A review of current leach operations solution chemistry will help to define acid consumption. This is recommended for all current leach operations.DISCUSSIONThe current practice at Freeport-McMoRan (FCX) is the use of column test results to define operating parameters. The monitoring of operational solution chemistry and copper production addresses any changes needed in operational parameters. This practice maximizes the value extracted from each tonne of ore leached."

Jan 1, 2008

By A. M. Ramadan

The composition of phosphate rocks, mainly used in manufacturing fertilizers, varies among deposits; therefore, phosphate rocks from different sources are expected to behave differently in acidulation processes. This paper examines the effect of preconcentration of phosphate ore from El-Sebaiya East, Egypt, using flotation on uranium leaching. A high-grade ore sample (27.3% P2O5, 168 mg/kg U) supplied by the Al-Nasr Mining Company, was subjected to reverse single-stage flotation to separate carbonate from the phosphate sample. Subsequently, P2O5 and U concentrations were 32.2% (92.1% recovery) and 225 mg/kg, respectively. After the leach process, the U concentration in the liquor was 319 mg/L.

Jan 1, 2014