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By G. S. Hume

Dr. K. A. Clark: No one will doubt the statement of fact that, during drilling operations through the bituminous sand beds in the Mildred-Ruth Lakes area, the core-barrel was found to contain bitumen with minor amounts of sand instead of bituminous sand. But many will doubt the interpretation of this fact as meaning that beds of bitumen occur interstratified with the bituminous sand beds. They will prefer to hold faith in the laws of physics, which deny that sand will float on oil or that fluid under pressure will not flow away through a permeable medium, until it is proved that the appearance of the oil in the core-barrel was not due to circumstances surrounding the drilling of the bituminous sand beds. The implication of the interpretation that sand is floating on oil is clear and it is unacceptable. Bituminous sand is unconsolidated sand more' or less saturated with a viscous oil. It is as difficult to conceive of bituminous sand lying on top of a body of the viscous oil as it is to conceive of unconsolidated sand more or less saturated with water lying on top of a body of water. The only difference between the two concepts is that the oil is viscous whereas the water is fluid. Oil-soaked sand placed on viscous oil would, when released, sink through the oil slowly, whereas water-soaked sand would sink through the water quickly.

Jan 1, 1947

By R. J. Henry

This paper describing the mining practices at the Teck-Hughes mines at Kirkland Lake, Ontario, is one of a series being prepared by the United States Bureau of Mines on mining practices, methods, and costs in various mining districts. HISTORY Prospecting was begun in the neighborhood of Kirkland Lake in 1906, but staking did not take place until 1911. Gold-bearing veins were discovered by surface trenching in 1913 on the mining claims which now make up the Teck-Hughes Gold Mines (Ltd.). A small mill was erected in 1916 after the completion of some underground exploration. A transmission' line was built in 1917 to the camp to provide adequate power for a vigorous underground campaign. Also in 1917 the present management took control operations since 1916 have been continuous,' with a steady increase in production to date.

Jan 1, 1930

The effect of superficial gas rate distribution (Jg profile) on down-the-bank metallurgical response was tested in the rougher-scavenger bank (eight 10-m3 tank cells, 4 roughers and 4 scavengers) at Highland Valley Copper Mine's Cu/Mo separation circuit. Three Jg profiles were tested: (a) as found (i.e., typical practice), (b) increasing decreasing (i.e., increasing from cell 1 to 4 and decreasing from cell 5 to 8), and (c) decreasing-increasing (i.e., decreasing from cell 1 to 4 and increasing from cell 5 to 8). The Jg profile approach proved to be of considerable value. Profile (b) gave the best metallurgical results, i.e., the highest down-the-bank grade/recovery (increase in grade at constant recovery was approx. 3-4% compared to the as found case) and the most efficient Cu/Mo separation (cumulative Cu recovery was reduced by almost half - approx. from 20% to 12%, at 99% Mo recovery). The success of the evaluation prompted the implementation of profile (b). Detailed gas dispersion characterization tests (Jg and Db (bubble size)) showed significant discrepancies between the Jg measured directly (using the McGill University sensor) and the Jg calculated from field instrument readings (volumetric gas flowmeter) and machine dimensions (cell cross-sectional area). Tests were designed to reveal the source of disagreement. It was observed that a significant number of bubbles were not able to float (i.e., they were entrained with the tailings), the evidence appearing to be a significant number of very small bubbles (approx. 200-300 µm) which did not have sufficient buoyancy to rise. Consequently, rather than relying on the field instrument, the evaluation depended on the McGill University Jg sensor, which collects bubbles in a tube near the pulp-froth interface, thus providing significance to the data as the measurements were obtained from bubbles that almost certainly floated.

Jan 1, 2007

"Beneficiation of low grade ores is the process of increasing the grade of a mineral through unit operations that remove and separate the waste (or gangue) material from the valuable mineral that can then be used for further processing or direct use. Important factors when deciding upon a beneficiation program are the feed grade, mass yield and the metal recovery. As lower grade ores are being processed beneficiation is a critical process. How predictive is laboratory test work for “as built” operating projects? Beneficiation includes all or some of the available unit operations to produce a beneficiation flow sheet i.e. associated unit operations including but not limited to optical sorting, radiometric sorting, crushing and screening, coarse cobbing, jigging, dense media separation (DMS), spirals, magnetic separation, flotation, thickening, filtering and drying. Reference to specific projects and the approach developed in formulating a beneficiation flow sheet and a comparison of difficult and simple beneficiation flow sheets are cited. INTRODUCTIONReserves of high-grade iron ores are reducing all over the world, at an alarming rate, over the years as a result of the rapid increase in demand for steel particularly from Japan and then China. The recovery of mineral value from the low-grade ores using present technology is more expensive than DSO ores due to high energy and capital costs. Beneficiation is defined as a variety of processes whereby extracted ore from mining is reduced to particles that can be separated into mineral and waste.Various physical properties of an ore can be exploited to separate a low grade fraction including variation in hardness, density, reflectance, conductivity and magnetism.""Beneficiation"" of iron ores includes concentration, generally by physical removal of unwanted gangue. Also considered beneficiation is the regulation of product size, or other steps such as agglomeration to improve its chemical or physical characteristics prior to smelting."

Jan 1, 2012

By Danqing Xu, A. W. Khair

"This paper presents a laboratory simulation of regrinding and resulting dust generation. An Automated Rotary Coal Cutting Simulator (ARCCS) was modified to conduct regrinding experiment. In order to estimate the effects of coal properties and operating parameters on regrinding and dust generation, an orthogonal fractional factorial experimental design was implemented in regrinding experiment. Marple personal cascade impactors, Series 290, were used to measure dust concentration during experiment. Three types of coal seams, namely, Waynesburg, Pittsburgh, and Lower Kittanning, were used to represent different coal properties. Experimental results and complete parametric analysis were presented.In a rotary cutting system, all the broken coal and part of the remaining coal from previous cutting cycles must be transported through part or entire circumference of the cutting drum. This transportation of already recovered coal through the cutting zone of rotary head makes it a rotary grinder. Some of the coal particles will be subjected to secondary fragmentation or regrinding in the process. The potential effect on dust generation by regrinding is significant because the cutting head utilizes kinetic energy to comminute recov¬ered coal material. The moving bits and bit blocks on the cutting head impact coal particles and transfer some of their kinetic energy to the particles. After obtaining kinetic energy, coal particles are free to impact one another and they also impact the coal seam or cutting head. During impact, high impact force and high internal stress concentration are created in the particles. Kinetic energy is transformed into strain energy. When stored strain energy exceeds allowable strain energy, the particles will break. A great amount of respirable dust will be generated due to impact failure. In reality, regrinding occurs at the same time as primary cutting. The particles to be ground are the result of primary cutting. Both primary cutting and regrinding produce dust. Therefore, it is impossible to separate regrinding from primary cutting. In order to study respirable dust generation by regrinding and analyze the influencing parameters, a laboratory approach was adopted. In this approach, it is possible to study the regrinding separating from primary cutting."

Mar 1, 1992

By T. P. Johnson, A. Vathavooran

"Conventional mining operations can produce a wide range of wastes containing gold and other precious metals but which are less suitable for processing by traditional methods. There are also types of ore that do not fit easily into a traditional processing flowsheet, e.g. because the refractory nature of their minerology or the combination of valuable metals is less suitable for traditional processes. Nevertheless, these potential sources of valuable metals remain unexploited resources that have attracted growing interest in the metals and mining industries over recent years. The challenge is therefore to develop and use alternative processing methods that have the flexibility of operation to cope with a wide variety of material types and tonnages, whilst providing high recoveries for valuable metals. Tetronics’ DC plasma smelting technology achieves industry-leading recovery rates for precious metals from a wide range of input materials and initial precious metal concentrations, whilst generating an inert slag-based material as a by-product (even when starting from a hazardous waste input material), approved within the EU for use in construction applications. This paper describes how this technology can be used as a key part of an overall process for the recovery of precious metals from these interesting materials.INTRODUCTION The extraction of gold from its ores by various forms of cyanide leaching has been practised by the mining industry for over a century (Minerals Council of Australia, 2005) and it remains the primary extraction method in a wide range of mining operations. The ability of cyanide to extract gold efficiently and quickly from lower grade ores without extensive additional processing, coupled with well-known methods to recycle the cyanide for re-use within the process mean it is often both the most economically preferable and environmentally acceptable method for recovering gold. As a result, 88% of all gold produced from mines in the USA in 2014 was recovered via leaching of open heaps or dumps (UGGS, 2016). Nevertheless, there are occasions where certain combinations of mineralogy and other circumstances reduce the effectiveness of cyanidation techniques, so that ever more elaborate additional processing steps are required, with inevitable increases in operating costs (Swash, 1988). In these circumstances, which are likely to be more common in future as ore qualities diminish, other methods of extraction may offer better alternatives. Whilst much less commonly used for Au extraction at present, pyrometallurgical methods are amongst the longest-established of these alternative approaches and merit serious consideration in such cases."

Jan 1, 2017

By J. O. Marsden

The Candelaria copper-gold mine is located approximately 25 km south and east of Copiapo in Region III, Chile. The original concentrator was designed to treat 28,000 t/d (30,800 stpd) of chalcopyrite-bearing ore through a circuit consisting of a single 11.0 m (36 ft) diameter SAG mill, two ball mills, conventional rougher and scavenger flotation, column cell cleaner flotation, and ceramic disc filtration. The plant was commissioned in August 1994 and quickly reached capacity and expected levels of metallurgical performance. In May 1996, a decision was made to double the capacity of the concentrator by the addition of a second line of equipment. The expanded facilities were commissioned in October 1997, less than 18 months after the decision to proceed, and well within the budget. The original concentrator facilities were replicated or duplicated as much as possible, however, a number of important equipment changes were made to take advantage of industry developments since the original concentrator was designed. These included the use of Eimco 120 m3 (4,200 ft3) cells for rougher and scavenger flotation and the installation of a Svedala VTM -800 Vertimill for regrinding. This paper describes the rationale behind these changes and explores the reasons for the success of the project.

Jan 1, 1999

By D. McElhinney, J. Winfield, G. Buchan, B. Slaker, A. Iannacchione, T. Minoski, G. Rashed, M. Van Dyke, M. Murphy

"The National Institute for Occupational Safety and Health (NIOSH) has previously established pillar design guidelines for the underground stone mining industry. These guidelines were created from an empirical database of pillar observations and were largely drawn from shallow, flat-lying mining operations. Current trends forecast a great amount of underground stone mines developing under deeper cover, higher seam dips, and more frequent multiple-seam mining configurations than in the past. The complex loading conditions and limited experience in these environments presents a substantial increased risk of ground failure. To investigate these loading conditions, NIOSH has installed a pillar and roof monitoring system at an underground limestone mine in central Pennsylvania. The mine is operating in a seam with dips ranging from 13 to 18 degrees and overburdens ranging from 275 m (900 ft) to 425 m (1,400 ft). Instrumentation installed in an undeveloped pillar and in surrounding strata is currently measuring pillar and roof behavior in response to stress redistribution during excavation. Laser scans are being used to measure ground displacement and changes in conditions associated with local mining conditions and significant geological features. Initial stress measurements show a gradual increase in stress in response to mining, and stress changes that correspond with blasts. The seismic system is operational, with calibration events signaling on each of the 18 sensors. INTRODUCTION Improving mine-wide ground control stability can help reduce the potential for catastrophic failure of underground structures. The National Institute for Occupational Safety and Health (NIOSH) has conducted prior research in stone pillar design and has established fundamental pillar design guidelines for the underground stone mining industry, which resulted in the development of the S-Pillar program (Esterhuizen et al., 2011; Esterhuizen and Murphy, 2015). The data that constitute the S-Pillar program were derived from past mining experience that was almost exclusively drawn from shallow, flat-lying, single-level mining operations1. Stakeholder interactions with NIOSH researchers have indicated that increased numbers of underground stone operations are expected to mine under significantly deeper cover, with greater seam dips, and under multi-level operating conditions (Slaker and Murphy, 2016). The complex loading conditions and limited experience in these challenging environments present a substantial increased risk of ground failure. Inadequate pillar design in these new and challenging environments could result in catastrophic wide-area pillar collapses and an increased safety burden to miners working in the U.S. limestone mining industry. To develop design guidelines and prevent inadequate pillar designs for these challenging environments, NIOSH is currently conducting detailed pillar response investigations at case-study mine sites operating in challenging environments."

Jan 1, 2018

By D. E. Cameron

"INTRODUCTION Grade control requirements for base and precious metal projects are an important input to both conceptual and feasibility mine plans. It’s the author’s experience that grade control input is insufficient at these critical stages. The few exceptions are a stimulus for this paper. A significant portion of CRC’s practice has involved assisting operations to solve grade control issues which were finally recognized as contributing to negative production variances. Even in operations perceived to be successful, deficiencies in grade control are often masked by measures taken by an operation to compensate a deficiency in one area with outperformance in another. The advantages in having a realistic and adequate grade control plan developed at an early project stage by a geologist(s) with considerable operations experience are manifold. It’s important for the geologist to prepare clear inputs into the mine plan backed by detailed supporting information in usable, flexible formats. When the decision is made to advance a project, the exploration team passes its model to a mine planner. The team doesn’t usually provide a detailed grade control plan with the block model, thus the planner tends to treat grade control as an activity to be rolled into a broader operating category. The consequences of highly compartmentalized mine planning can be severe in the long run. Once a plan is tabled, and especially once production begins, it is very difficult to improve grade control outside the fixed parameters of the operation. Major fixes tend to be costly in terms of time and dollars. Implementing needed changes to the grade control practices and culture in an established mine can present almost insurmountable difficulties. Hypothetical examples of grade control planning are presented to illustrate the author’s points. The diagrams are conceptual and schematic; any resemblance to actual deposits is coincidental. BEYOND 50¢/TON In more than a few cases, grade control cost is buried in a subcategory of mining cost with little more consideration than a plugged number. It is possible to do a word search in 43-101 Technical Reports on advanced projects and without finding reference to “ore control” or “grade control”. The idea is to reach the orebody by the shortest and least expensive means, achieving production and investor payback at the earliest possible occasion. This entails a low stripping ratio pit or a decline ramp with sills on ore to prepare the first stopes. The processes to achieve the target production, one of which is grade control, will fit into this framework. After all, there is a block model and the planner can see the grade and tons in each block. The exploration geologist whose responsibility was to bring a large percentage of the deposit into mineral resource owns the model and has released it to planning. Now comes the planning stage. Mine geologists haven’t been hired yet and they will be phased in with production. What could go wrong? Larger companies may have the geologists with operating depth who will be inserted into the mine planning discussions, but it’s not a given. Junior companies must often delegate the conceptual and detailed planning tasks to consultant groups. Is a senior-level grade control review part of the scope of work for the mine planning? If a good grade control plan is in place before the mine geology team is staffed, a key element of success is assured."

Jan 1, 2018

By K. Yamaguchi

To contribute highly efficient and environmentally conscious copper smelting process, several topics in chemistry of the continuous converting were discussed. Through the reviewing the authors' thermodynamic simulations, it is recognized that small amount and low copper content of the coexisting slag is preferable. Some discussions were added to the Takeda's diagram for Cu20-FeOx-CaO ferrite slag, and the activity curves of the constituents were derived. Thermodynamic calculations were carried out on the copper contents of the slag during the blister copper formation. The dissolution of copper in slag decrease somewhat with increasing temperature, but the effects of increasing S02 pressure and decreasing %S in copper are considerable to increase in copper content of the slag. The optimum oxygen enrichment in the supplied air, and regulation of the oxidation degree for the blister copper are important in practice. The gypsum phase separation in the calcium ferrite slag was also evaluated thermodynamically, and concluded that decreasing temperature, increasing CaO content or S02 pressure, and extensive oxidation of the blister copper tend to separate CaS04.

Jan 1, 2007

By D. Milne, G. Snell

"A new method to predict stope hangingwall failure in underground excavations is proposed. This method compares movement rate for stope hangingwalls, determined from changes in strain or deformation, with changing opening geometry. This geometry is represented by the effective radius factor (ERF), a measure of relative distance to supporting abutments from an opening surface point. This strain-ERF (SERF) approach is applied to an instrumentation program for Brunswick mine near Bathurst, New Brunswick. The SERF method for this case study predicted failure at an ERF of 13.9 m and hangingwall failure actually occurred at an approximate ERF of 12.4 m.RÉSUMÉ Cet article propose une nouvelle méthode permettant de prévoir la rupture de l’éponte supérieure des chambres d’abattage dans les excavations souterraines. Cette méthode compare la vitesse de mouvement des épontes supérieures des chambres d’abattage, déterminée à partir des changements aux niveaux des contraintes ou de la déformation, à la géométrie variable de l’ouverture au jour. Cette géométrie est représentée par le facteur du rayon effectif (ERF, de l’anglais Effective Radius Factor), une mesure de la distance relative jusqu’aux appuis latéraux de soutènement à partir d’un point de la surface d’ouverture au jour. Cette approche contraintes-ERF (SERF, de l’anglais strain-ERF) est appliquée à un programme d’instrumentation pour la mine Brunswick située près de Bathurst, dans le Nouveau-Brunswick. Pour cette étude de cas, la méthode SERF prévoyait une rupture à un ERF de 13,9 mètres (m); la rupture de l’éponte supérieure s’est effectivement produite à un ERF d’environ 12,4 m.INTRODUCTIONUnderground stope design can be improved with the installation of field instrumentation. The most commonly used instrumentation for monitoring rock mass behaviour around open stopes is extensometers. The y-axis of Figure 1 represents how the changing rate of movement, measured using either displacement- or stress-type instruments, can be observed over time. The magnitude of movement measured by extensometers can give an approximation of the load on installed support (Milne, Pakalnis, & Lunder, 1996) and a changing rate of movement can indicate approaching instability."

Jan 1, 2018

By Anthony (Sam) J. S. Spearing

Backfilling using coal waste is not a novel means for waste disposal throughout the world; yet, backfilling using a high density (paste) backfill is relatively new, except in Germany and Poland. One of the main benefits of a high density backfill, as opposed to a low density (slurry) backfill, is that, if placed at a low permeability and low moisture content, the risk of water table contamination can be relatively low. In addition, if paste backfill can be placed cost effectively, environmental benefits are huge because the surface impoundment footprint can be greatly reduced or even eliminated. This is especially true in room-and-pillar applications. In its current state, backfilling is considerably more costly than surface disposal. This is mainly because paste, as opposed to slurry, significantly increases backfill placement costs; however, this could be offset by increasing the achievable extraction rate. Overall, such a system could potentially improve the public?s image of coal mining in the US for a myriad of reasons. This paper discusses a proposed backfill system for a typical room-and-pillar coal mine in the Illinois Basin of the US and clearly shows the technical advantages, although costs are still higher with fill. The cost differential will likely reduce over time, in favor of the backfill option, as costs of traditional surface disposal increase, due mainly to ever-restrictive disposal regulations.

Jan 1, 2013

Is your company paying taxation on profits which are partly illusory?ò Are unions making wage demands based on erroneous notions of the real level of value added to your companyÆs production?ò Do you receive historic data on profit contributions by different products, or for different market areas, which you feel may be misstated?ò Do you find that cash is not always readily available to pay dividends out of reported earnings because, it is required to finance additional working capital without a real increase in the level of activity?ò Are your planning and budgeting routines hampered by the need to rely on historic financial data which bear little resemblance to what the future may hold for your company?ò Are you able to maintain your business at its optimum operating capability or do you feel each year that it becomes more and more difficult to achieve the same volumes of prdduction at the same level of return on assets employed without having to raise more capital?If your answer to any of these questions is "yes", then it is very likely that Current Cost Accounting (CCA) has something to offer you!characteristics of the mining industry and for single venture mining operations.

Jan 1, 1984

RISKGATE is an interactive online risk management system developed by the Minerals Industry Safety and Health Centre (MISHC; University of Queensland) to maximise health and safety performance in the mining industry through implementing continual improvement in management of major unwanted events. A comprehensive knowledge base has been captured for risk management of tyres, collisions, fires, isolation, strata underground, ground control open cut, explosions, explosives, manual tasks and slips/trips/falls. This will be extended to outburst, coal burst and bumps, interface displays and controls, tailings dams and inrush in 2013. RISKGATE presents substantive and leading edge controls to assist industry stakeholders in the design, management and reporting of organisational and regulatory compliance requirements. From a broader industry perspective, RISKGATE provides an environment for knowledge capture and knowledge exchange to drive innovation and best practice in the identification, assessment and management of risk. By capturing operational knowledge from industry experts, RISKGATE provides a cumulative corporate memory at a time of high personnel turnover in the coal industry. All major Australian mining companies are contributing topic experts to this broad industry initiative. In the first 24 months, nine coal mining companies nominated over 80 different experts who contributed over 400 days of individual time to the knowledge capture workshops. RISKGATE is the largest single Australian Coal Association Research Program (ACARP) occupational health and safety (OHS) initiative to date.

Aug 1, 2013

By A. Pomianowski

A spectrophotometric study of xanthate solutions alone and their mixtures with equimolar amounts of alkyl trimethyl ammonium bromide (CnTAB) has established the additivity of ultraviolet spectra (190 - 400 my) for potassium ethyl xanthate (KEtX) and C12TAB. No obvious effect of C12TAB addition on the decomposition rate of KEtX was observed. The kinetics of collector abstraction by copper, nickel and sphalerite powders were studied in 10-5 - 10-3M solutions of KEtX alone and in equimolar mixtures of KEtX and C12TAB. Presence of C12TAB caused a 2 - 3 fold increase in the amounts of xanthate (equivalent to many monolayers) abstracted by copper and, nickel powders but had no effect on the abstraction of xanthate by unactivated sphalerite. The adsorption of C12TAB itself from the mixed solutions was very low, at most ore monolayer only, and was not influenced by the nature of the solid powder used. The results are interpreted in terms of molecular associations and synergistic action of xanthates and quaternary amines.

Jan 1, 1963

By A. Anani, K. Awuah-Offei

"Accounting for changing duty cycles in equipment matching maximizes equipment utilization and productivity. The objective of this study is to investigate if the optimal number of shuttle cars for a particular panel width is optimal in different segments of the panel. In this study, discrete event simulation (DES) is used to determine the optimal number of cars that maximizes productivity in each segment. Data from a real-life mine is used to validate the model. The study shows that the optimal number of cars for the entire panel may not be optimal for some of the defined segments. INTRODUCTION Fleet optimization is important for efficient mine management. Deploying the optimal fleet minimizes the capital and operating expenditure for particular operating conditions and production targets. In an operating mine, with an already existing fleet, deploying the optimal fleet maximizes productivity or minimizes production cost with the available resources. Hence, a lot of research has focused on how to determine the optimal fleet size for a mining operation, given particular operating conditions. Mine managers and engineers have successfully used discrete event simulation (DES) to determine the optimal fleet size in mines (Michalakopoulos, Roumpos, Galetakis, & Panagiotou, 2015; Awuah-Offei, Osei, & Askari-Nasab, 2011; Ben-Awuah, Kalantari, Pourrahimian, & Askari-Nasab, 2010; Govinda Raj, Vardhan, & Rao, 2009; Yuriy & Vayenas, 2008; Kwame Awuah-Offei, Temeng, & Al-Hassan, 2003). In most cases, the published work on this topic involves determining the number of hauling units to be matched to a loading unit. Many of these examples involve truck-shovel optimization. There are, however, some underground mining examples. For instance, we recently concluded work that involves fleet optimization in underground room and pillar (R&P) coal mining (Awuah-Offei & Anani, 2015). Generally, when DES is used in fleet optimization it involves three steps: (1) Build a valid model of the mining system capable of predicting the required outputs with varying fleet composition; (2) Design and conduct experiments that vary the fleet composition across all feasible options; and (3) Determine the optimal fleet based on the results of the simulation experiments. A major decision in the first step is which outputs should be modeled. Some examples in the literature include productivity, unit cost, loader utilization, queue length, and hauler waiting times (Awuah-Offei et al., 2011; Govinda Raj et al., 2009; Kwame Awuah-Offei et al., 2003). Often, more than one of these outputs are simulated and considered as the objective function of the optimization."

Jan 1, 2016

By J. A. Herbst

It is widely recognized that Semiautogenous Grinding (SAG) circuits require the use of effective process control techniques to achieve their full processing potential and to protect them from the harmful effects of overloads and grind-outs. This requirement is even more essential as the variability of the ore being processed increases. The new Chino Mines Company concentrator located near Santa Rita, New Mexico started operation in June of 1982. At least eight major rock types are mined and milled. Ore hardness, feed size, and specific gravity are quite variable. Typical feed rates per grinding line are 850-900 tph but extremes of 260-1200 tph have been encountered. Normal feed rate variations of between 400 and 1100 tph are encountered on a weekly and sometimes daily basis. Feed size varies with rock type from a maximum of eight inches for the harder ores to two inches for the softer ores. One somewhat unique characteristic of the Santa Rita ore body among major copper mines is the wide variation in ore specific gravity. Specific gravities of the major rock types range between 2.7 and 3.6. The specific gravity variations are due to large changes in magnetite concentration throughout the pit. The result of this is large variations in SAG mill throughput which can have a harmful effect on downstream ball milling and flotation performance. Due to these wide fluctuations in ore characteristics, it was recognized soon after startup that successful operation of the grinding circuit and the plant at large was dependent upon the development of an effective automatic grinding control strategy. Research at the University of Utah has suggested that control based upon critical but difficult or costly to measure variables would improve automatic control performance. A project team made up of the Chino process control staff, Kennecott's Process Control department and the University of Utah Metallurgy Department was established to assess these concepts at Chino. This paper describes the development and on-line plant implementation of an Extended Kalman Filter that uses simple models of the SAG milling process, to predict the volumetric loading, the rock mass, the water mass, and the ball mass within the SAG mills at Chino. The on-line estimation of these parameters is the first step in the development of an integrated SAG mill/ball mill control strategy which is hoped to L eventually maintain the grinding circuit at its maximum capacity, while balancing the grinding load between the SAG and ball mills. Examples of long term estimation with the models and filter are shown for typical operating conditions as well as a SAG mill overload and grind-out.

Jan 1, 1983

By Jerry E. Olson

In 1984, the U.S. Bureau of Mines conducted a mineral survey of the Black Rock Desert Wilderness Study Area (WSA) (NV-020-620), which covers approximately 319,000 acres in Humboldt County, Nevada. About 90% of the WSA is underlain by Pleistocene desert playa sediments; the remainder is comprised of Jurassic-Triassic metamorphic rocks, a Cretaceous intrusive and Tertiary volcanic rocks. Most of the WSA is within a down-dropped fault block of the Basin and Range physiographic province. No mines or evidence of major exploration were identified within the WSA; but five claim blocks within or adjacent to the WSA and zones of anomalous character as defined by the Bureau of Land Management were examined. No mineral resources were identified, but anomalous gold, silver, and indicator element content of samples taken near Pinto Hot Springs indicate resources may exist. Lithium content of samples from near-surface sediments on the desert playa is too low for resources, out geologic conditions are favorable for higher concentrations at depth. Occurrences of zeolites are not considered to be of commercial interest. Approximately 27,000 acres are held under geothermal lease, and most of the WSA is of interest for oil and gas. No evaluation of these commodities was made by the Bureau of Mines

Jan 1, 1985

By P. J. Petit

Changing mining conditions, legislative issues, rising costs, and the constrained supply of skills and electricity in South Africa, have been some of the criteria forcing mine operators and engineers to review the traditional use of compressed air to power underground hand-held drilling operations in narrow-stope, hard-rock mines. In answering the question, 'What is the best way to power this mine?', they have resorted to trade-off studies with results of limited usefulness, because they are site-specific, and use only a few comparative parameters, based on a small data set. This paper aims to provide definitive, objective and quantitative decision-making approaches to evaluate alternative energy delivery systems for stoping in narrow-reef hard rock mines, based on primary and relevant criteria. These include quantifiable criteria such as capital expenditure, operating costs, resulting financial returns, energy and water usage, production performance, life-of-mine, etc. Other important criteria with qualitative aspects are integrated. These are safety, health and environmental concerns, legislative requirements, business risk, technology change, etc. The paper begins with a review of decision making processes available, and proposes the use of the analytical hierarchy process (AHP) method as a multiple-criteria decision making tool and the decision tree-based on milestone-driven uncertainties, to validate the selection of the energy delivery system. The results of a multiple-criteria questionnaire completed by a variety of industry experts and professionals, such as production and operation managers, consultants, suppliers and techno-financial analysts, are discussed. The evaluation indicates the potential for hydropower to be the best solution for narrow reef hard rock mines, based on current information, mine design layout and production.

Jan 1, 2012

By K. Yamaguchi

To contribute highly efficient and environmentally conscious copper smelting process, several topics in chemistry of the continuous converting were discussed. Through the reviewing the authors' thermodynamic simulations, it is recognized that small amount and low copper content of the coexisting slag is preferable. Some discussions were added to the Takeda's diagram for Cu20-Fe0x-Ca0 ferrite slag, and the activity curves of the constituents were derived. Thermodynamic calculations were carried out on the copper contents of the slag during the blister copper formation. The dissolution of copper in slag decrease somewhat with increasing temperature, but the effects of increasing S02 pressure and decreasing %S in copper are considerable to increase in copper content of the slag. The optimum oxygen enrichment in the supplied air, and regulation of the oxidation degree for the blister copper are important in practice. The gypsum phase separation in the calcium ferrite slag was also evaluated thermodynamically, and concluded that decreasing temperature, increasing CaO content or S02 pressure, and extensive oxidation of the blister copper tend to separate CaS04. 0

Jan 1, 2007