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By J. B. Hadaway

The current limit to power transmission for grinding mills through one mill pinion is approximately 7,500kW. Above this power rating either twin pinion arrangements or Gearless drives are required. The SMTP concept was developed to overcome this limitation by using two pinions with a single large drive motor. The SMTP mill drive offers major advantages in capital cost and greatly simplifies plant layouts, while using an innovative adaptation of industry-proven single helical gearing. Through the use of a single motor, provision of variable speed drive is greatly simplified. The SMTP can also be used for significant power upgrades on existing mills.

Jan 1, 2001

By Z. T. Bieniawski, E. H. Skinner, R. A. Cummings, F. S. Kendorski

INTRODUCTION Caving Mine Drift Support Design The support of production drifts remains a high-cost item facing the developer or operator of an underground caving mine. Support costs are incurred well in advance of ore production, and there is a strong need to minimize these pre-production commitments of capital. Due to the extensive rock loads that normally develop during the caving process, the support used in caving mines is typically quite heavy and therefore costly. The process of designing this support increasingly makes use of basic tunneling principles, but is hindered by the still-incomplete state of knowledge of rock behavior in caving mines, and high variability of geological conditions in caving orebodies. Caving mines differ from tunnels in that the deformations and rock loads induced are transient and dynamic in occurrence, and are normally large in magnitude, often well in excess of rock mass strength. As opposed to tunnels, caving mine layouts are more intricate, and the methods and effects of construction ordinarily are not as closely monitored. Safety factors in mining tend to be lower and rock masses tend to he weaker. These aspects prevent strict application of tunnel design technology to caving mine drift support, with the result that mine drift supports are commonly designed chiefly on the basis of experience at other mines. Mining procedures at caving mines are evolved in response to site-specific requirements, to the extent that there are really as many caving methods as there are caving mines.

Jan 1, 1983

By S. C. Chelgani

The Thor lake deposit is a world class resource of rare earth (REE) metals and minerals in Canada. Development work to optimize a REE mineral recovery process flow sheet is underway, however, given the ore mineralogy; the developed reagent scheme is relatively complex. As part of a research project, micro flotation tests were conducted on a feed sample in order to examine factors affecting stream partitioning. SEM-EDX was performed to evaluate variability in grain composition between streams (concentrate and tails) and TOF-SIMS surface analysis was used to determine statistically significant differences in surface species particularly related to potential activation (or depression) of the examined mineral phases. SEM-EDX analysis reveal that the concentrate has a significantly higher proportion of REE bearing grains (carbonates and phosphates) relative to the tail (almost none were identified). Spectral fingerprinting by TOF-SIMS has allowed for the identification of all reagent species investigated. Reagent signal intensity discrimination on test stream mineral surfaces was observed by the TOF-SIMS analysis using reagents at plant concentration levels. TOF-SIMS analysis confirmed that REE bearing grains reporting to the concentrate are doing so in response to collector attachment whereas grains reporting to the tail are doing so in response to a lack of collector and/or in combination with the presence of the depressant. The surface analysis of gangue phases reveal a similar reagent discrimination; the signal intensity of collector species was significantly higher on the concentrate samples relative to the tails while depressant species were significantly enriched on the surface of the gangue phases in the tail samples. A detailed evaluation of the surface species representing the various reagents used in flotation scheme revealed a distinct competitive relationship between two of the reagents. The surface analysis identified that when used in concurrently, there appears to be a negative feedback resulting in a significant reduction in loading for several of the collectors on grains reporting to the concentrate. An evaluation of the effect of reagents on REE mineral in pilot plant is currently under way. Key words: Rare earth minerals, Flotation, Surface chemistry, TOF-SIMS, SEM-EDX

Feb 27, 2013

By K. Huttu, A. van Wageningen

"The Kittila underground mine is extracting one of the largest known gold deposits in Europe and has an estimated mine life through 2034. The Kittila mine is located in the Lapland region of northern Finland, approximately 900 kilometres north of Helsinki and 150 kilometres north of the Arctic Circle. The ore at the Kittila mine is refractory, making gold extraction relatively difficult because the gold is generally locked inside the two main sulphide minerals – arsenopyrite and arsenic-rich pyrite. The orebody consists of a complex pinch-and-swell structure including up to 100 different gold mineralization lenses. Fluctuating gold prices and other economic considerations increase the need to maintain and update the mining plans more frequently. Over 3,000 stopes need to be designed and optimized once a year for the LOM-plan and more frequent updates are needed for mid- and short term plans as well as production studies. As a large part of ore body is very sensitive to cost and gold price changes, stope optimization is crucial to find the best possible production scenario for every moment. Kittilä has gone through several methods for stope design and optimization. Most recently the Deswik Stope Optimizer (Deswik.SO) has been successfully used. Dewsik.SO has helped Kittilä to design and optimize stopes in a timely and cost efficient manner. Together with the Deswik suite of planning tools, reliable and auditable plans are created. This paper describes the methods used for stope design and optimization at Kittilä. THE KITTILÄ MINE The Kittila underground mine in northern Finland is extracting one of the largest known gold deposits in Europe. Kittila achieved commercial production on May 1, 2009, becoming Agnico Eagle’s first mine to open outside of Canada. The Kittilä orebodies were initially mined from two open pits – Suuri and Roura – and underground operations were added in October 2010. The pits were mined out in November 2012, so mining is now entirely underground in the Suuri and Roura deposits. Additional, small open pits will mine any remaining near-surface reserves in the future. The Kittila mine is located in the Lapland region of northern Finland, approximately 900 kilometres north of Helsinki and 150 kilometres north of the Arctic Circle. With a mine life estimated through 2034, Kittila is Agnico’s longest-life mine. The mine’s proven and probable reserves contain 4.7 million ounces (32 million tonnes at 4.6 g/t gold). Ore has been mined from underground since 2010. The original 3,000-tonne-per-day operation underwent an expansion project to increase the throughput capacity of the mill by 1,000 tonnes/day, starting by the end of 2014. The region around the Kittila mine is underlain by mafic volcanic and sedimentary rocks of the Kittila Greenstone Belt oriented nearly vertical. This greenstone belt is similar to those hosting Agnico Eagle's Canadian deposits in Quebec’s Abitibi region and Nunavut. At the contact point between iron-rich and magnesium-rich rocks lies a 100- to 200-metre-thick structural zone known as the “Suurikuusikko Trend”. This trend hosts the Kittila deposit, where multiple mineralized zones have been traced over a strike length of more than 25 kilometres."

Jan 1, 2015

By R. G. Benson, P. R. Wotruba, K. W. Schmidt

Introduction The Fortitude deposit is located in Nevada's Battle Mountain mining district in the central Great Basin region of the Basin and Range province of the western US. The deposit is at Copper Canyon in the southernmost portion of the district, about 19 km (12 miles) southwest of Battle Mountain (Fig. 1). Mining activity in the district began in 1866 (Roberts and Arnold, 1965). Base and precious metals were mined by underground methods intermittently until 1955. Placer gold was first discovered at Copper Canyon in 1912. Moderate scale dredging of alluvial gravel at the mouth of Copper Canyon was conducted by the Natomas Co. from 1944 to 1955 (Johnson, 1973). In 1967, Duval Corp. began large scale open-pit copper mining. (Sayers, Tippett, and Fields, 1968). Copper prices decreased and precious metals prices increased in the mid-1970s so Duval's exploration efforts at Copper Canyon shifted to precious metals. These exploration efforts resulted in the discovery of the Tomboy-Minnie deposits and shortly afterward, the Northeast Extension deposit at Copper Canyon (Blake, Wotruba, and Theodore, 1984; Fig. 2). The Fortitude deposit was discovered in late 1980, after Duval had proceeded with precious metals production at Copper Canyon. Converting the existing plant facilities at Copper Canyon to a precious metals recovery system was completed in 1978 (Jackson, 1982). Production from the Tomboy-Minnie deposits began in early 1979. Precious metals production moved to the Northeast Extension deposit in 1982. In late 1984, after three years of preproduction stripping and mining of the Upper Fortitude ore zone, production from the higher grade Lower Fortitude ore zone began. Currently, Battle Mountain Gold Co. operates the Fortitude deposit. Production from the Battle Mountain mining district before 1967 included 4.6 t (150,000 oz) of gold, 65.3 t (2.1 million oz) of silver, 13.6 kt (15,000 st) of copper, 4.5 kt (5,000 st) of lead, and 13.6 kt (15,000 st) of zinc (Roberts and Arnold, 1965). At the time of Duval's 1967 start-up, ore reserves at Copper Canyon included 12.6 Mt (14 million st) in the East ore body and another 3.6 Mt (4 million st) in the West ore body. Both had average grades of 0.79% copper, 16.1 g/t (0.47 oz per st) silver and 0.86 g/t (0.025 oz per st) gold (Blake et al., 1984). The Tomboy-Minnie deposits collectively contained 3.5 Mt (3.9 million st) of ore grading 3.09 g/t (0.09 oz per st) gold and 9.6 g/t (0.28 oz per st) silver. The Northeast Extension reserves were identified as 1.4 Mt (1.5 million st) of ore with an average grade of 2.9 g/t (0.085 oz per st) gold and 5.1 g/t (0.15 oz per st) silver. Initial total reserve estimates for the Fortitude deposit were 14.5 Mt (16 million st) grading 5.14 g/t (0.15 oz per st) gold and 29.8 g/t (0.87 oz per st) silver (Jackson, 1982). Therefore, before large scale mining, the Copper Canyon area contained a minimum of 102.6 t (3.3 million oz) of gold and 541 t (17.4 million oz) of silver as mineable reserves. The Fortitude deposit reserves as of Jan. 1, 1985 were 10.6 Mt (11 million st) grading at 5.28 g/t (0.154 oz per st) gold and 28.5 g/t (0.83 oz per st) silver. Currently, the company is producing 6.8 t (220,000 oz) of gold and significant amounts of

Jan 7, 1986

The National Institute for Occupational Safety and Health (NIOSH) conducted a monitoring field study to optimize the rib control plan in a room-and-pillar coal mine. The study mine is located in the Central Appalachia coal basin. Two monitoring sites were selected at different sections in the study mine. The selected sites have different overburden depths and geologies. The instrumentation plans at study sites include borehole pressure cells installed at various depths in studied pillars, rib and roof extensometers and load cells mounted on number of the rib bolts. The instrumentation results and observed geology for one of the study sites showed great potential for optimizing the rib bolt plan at that section of the mine. Therefore, a rib support optimization study was followed using numerical modeling technique. A calibrated threedimensional model was developed using the instrumentation results of the optimization section. The calibrated model was used to determine the optimum rib support for expected overburden depths and rib heights. Baseline rib bolt patterns were assumed for the optimization study. Additional rib bolts, above and beyond the baseline bolting parameters, were outlined and identified as solutions to contain the rib fracturing identified by the modeling results.

Jul 28, 2020

By A. M. Elmahdy

An amphoteric collector was tested for dolomite removal from phosphate ore. The effects of some operating parameters and their significance on the flotation of dolomite were studied using statistically designed experiments. The parameters studied were impeller speed, pH and collector dosage. Analysis of the results showed that pH is the most significant parameter. Moreover, a concentrate containing 0.53% MgO and ~31.1% P2O5 with a recovery of ~ 90% can be achieved at optimum values of studied parameters, i.e., a pH of 12, an impeller speed of 1,000 rpm and a collector dosage of 2.5 kg/t.

Jan 1, 2007

By John B. Sabey, Vernon R. Degner

The industrial coal processing trend toward greater fines production, due to mining techniques and customer requirements for reduced ash and sulphur content in the fine coal product, is the basis for ongoing efforts to improve the performance of the flotation machine. WCO has successfully adapted the principles of the beds Column to a full size machine prototype. Recent plant evaluation tests confirmed the initial promise of this concept by achieving a considerably lower ash product while maintaining the combustible recovery level of conventional flotation machinery. The paper will include a review of the fundamental principles of the beds Column concept, together with a summary of the WEMCO R&D Program from concept evaluation and laboratory testing, through pilot, to full size machine tests. The performance of the WEMCO/Leeds flotation column will be compared to conventional flotation concepts using a flotation "kinetics" approach, which quantifies the floatable and non-floatable specie recovery performance using the Klimpel model as the basis for analysis. The variety of distinct features unique to this concept include a series of barrier rods which are used to advantage, often together with countercurrent reflux water flow, to achieve a significant improvement in product grade. The hydrodynamics associated with the mechanics of "gangue stripping" from the rising "air bubble-floatable specie" matrix will be briefly reviewed.

Jan 1, 1988

By K. L. Boyd

"Can the ever-increasing tonnages of ore and waste being mined today be accommodated simply by using bigger and better crushers? This paper reviews the latest primary crushing and large cone crusher designs. It highlights how—in addition to using the larger crushers now available—maximizing the utilization percentages in each of the four stages of the rock size reduction circuit will assist in meeting the larger production requirements of existing and new mining operations. Its goal is to help consultants and operators select the optimal equipment and custom-designed items in the material handling flowsheet from the mine to the mill, maximizing operating hours per year. INTRODUCTION Are larger crushers the answer to meeting the demands of existing or new mining projects for increased tonnage? Boyd (2002) predicted that larger crushers for primary and fine crushing would be beneficial and contribute to the further development of mineral processing. Larger crushers for primary and fine crushing applications are now available. The rock size reduction process from the mine to the selected milling process facility has not changed since man first broke a rock to get at the interesting mineralization within. The rock size reduction process circuit from the mine to the mill can be summarized in four stages: • Stage 1: Mining (rock breakage and fragmentation) • Stage 2: Pre-screening (scalping) • Stage 3: Primary crushing • Stage 4: Fine crushing (secondary, tertiary and quaternary) Plant throughput and ore delivery schedule will form the basis for flowsheet design and equipment selection. Once the project design criteria and the mine plan have been established, the circuit’s utilization percentage (percentage of hours when the circuit is being utilized) can be optimized for each of the four stages. The mining and process departments are the key stakeholders in the selection of the final comminution circuit, and both departments should be familiar with and contribute to the design of the circuit. This paper focuses on the four stages of a rock size reduction circuit from the mine to the mill, equipment selection, various ore and waste handling flowsheets, and improvements in crusher station design, all with the purpose of maximizing availability and plant utilization percentages. What is a large tonnage? From 1907 to 1914, the American Panama Canal excavated 300 million tonnes of material in 7 years (History of the Panama Canal, Wikipedia). Today’s large mining operations can move the same material in 10 months, using large crushers or multiple crushers for ore processing and large crushers and a mixture of truck haulage and conveyors for handling the large waste tonnages. Some examples include:"

Jan 1, 2017

By C. Bazin

The separation of rare earth elements is a complex process that is industrially carried out using liquid/liquid extraction. Detailed flow sheets of solvent extraction plants for the separation of rare earth elements are rarely available in the literature which limits the opportunity to teach and test the application of data reconciliation, process optimization and control to these types of plants. The basis of a simulator for a rare earth elements separation plant is described in this paper. The simulator is used to adjust the organic phase and stripping acid flow rate in a separation circuit of heavy rare earth elements from light one?s.

Feb 27, 2013

By António Fiúza, Anthony Danko, Patrícia Leitão, Aurora Futuro, Cristina Vila, Lurdes Dinis

The average global recovery rates for processing scheelite ores by physical methods are normally in the range 50 to 70%, and it is problematic to attain a minimum grade of 60% WO3 in the concentrates without penalizing further this yield. Direct hydrometallurgical processing of low-grade concentrates or medium- to high-grade ores may become an interesting global economic alternative as it allows increasing the recovery rate, avoiding major penalizations due to the grade of the concentrate and increasing the added value of the final product. Leaching of scheelite by sodium carbonate presents several advantages such as its selectivity regarding the dissolution of troubling metals like iron and aluminium, allowing for a simple and flexible process diagram. However, it has the disadvantage that it is performed at high temperature and pressure with a high consumption of reagents. The feasibility of leaching is appraised by the extraction yield and selectivity, but also by the operating cost. Considering these aspects, a relevant objective of the research was to transform alkaline pressure leaching into a feasible processing alternative directly applicable to both medium- and high-grade ores and simultaneously to low- and high-grade concentrates produced by physical methods. The research used ore samples and low-grade concentrates from a typical skarn orebody as well as high-grade concentrates from a chalcopyrite-scheelite orebody. Different operating conditions were tested. Temperature, pressure and reagent concentration showed to be fundamental, but not exclusive, in determining the leaching behaviour. The operating conditions are milder than the ones generally proposed in the literature or currently in use in industry, but with higher carbonate consumptions. The research was oriented towards potential industrial implementation allowing for the application of optimization methods and straight scale-up to pilot-plan design and project.

By S. Lolon

Spontaneous combustion in underground coal mines has become a serious problem particularly in the caved area (gob). Recent statistic has shown that approximately 17% of total 87 underground coal mine fires in the United States are attributed to spontaneous combustion (De Rosa, 2004). Spontaneous combustion results from a self-heating process in exothermic conditions. The accumulated heat, if not removed, is conducive to the rapid increase of temperature and may result in mine fires or explosions. It is well accepted that the interaction between oxygen and coal substances is the main cause for spontaneous combustion, while other factors such as pyrite, moisture, and bacteria play a secondary role to the self-heating of coal. Therefore, only coal oxidation is considered in this study. Coal oxidation occurs as coal comes into contact with air. This process involves complex phenomena in terms of heat transfer, chemical surface absorption, and energy balance related to inherent properties of coal (Wang et al., 2003). For simulation purpose, overall reactions can be simplified as suggested by Mitchell (1990): C + O2 CO2 + heat [65°- 94°C] (1) CO2 + C 2CO + heat [100°- 150°C] (2) Stoichiometric reactions (1) and (2) show that 2.66 grams of oxygen are required to oxidize 1.0 gram of carbon. Once the temperature exceeds 100°C, the reaction has little chance to stop. The crucial step in reducing spontaneous combustion risk is locating the ignition point of spontaneous combustion (hot spot). This information is useful in the effort of developing a preventive method effectively. In this study, this was conducted based on the best gathered information: ventilation surveys conducted in an existing longwall mine located in the western US; the laboratory experiments performed on a physical gob model; and the CFD models to investigate the flow behavior in the gob, the oxidation of coal, and heat transfer phenomena. The hot spot location is determined as a function of oxygen concentration and gob temperature. The critical values are the following: 5% (by volume) for oxygen and 100°C for gob temperature. CHARACTERISTICS OF GOB MATERIAL AND MODELS Permeability Concept Permeability is one of the key parameters in any study dealing with porous medium. This is determined by particle size and the ratio of void volume to the total volume (porosity) of porous medium. This relationship is known as the Carman-Kozeny equation (Scheidegger, 1957) and is given mathematically as [ ] where k* = theoretical specific permeability, m2 dm = the mean particle size, m n = porosity However, permeability tests with various particle sizes were carried out at the University of Utah?s ventilation model (Figure 1) to verify this relationship. These tests indicated a necessity to modify this theoretical relationship. A factor of 0.898 was obtained to make up the different between theoretical permeability and the experimental one. The modified version of (3) is given by: [ ] This equation is used to determine permeability of the simulated mine gob throughout the study. [ ] Mine Gob Material The gob is represented by zones filled with material of given size distribution. For physical measurements and CFD simulations, the material used were crushed rock and coal. Due to compaction, three different characteristics of gob material were used in the model: unconsolidated, semiconsolidated, and consolidated. In the real condition, the largest coal-rock particles are more likely to be located in the area behind the shields. This material is freshly broken and unconsolidated. The size of these broken particles is based on a study carried out by analyzing images taken from the area behind the shields in three coal mines in the United States (Pappas and Mark, 1993). The results have shown that the mean size of materials behind the shields is about 1.22 m. The permeability associated with this size was then obtained from (4) and permeability tests. This was 4.68 x10-7 m2. For semiconsolidated and consolidated zones, due to the lack of experimental data, these were determined through CFD simulations. These simulations were designed by assigning a permeability of 3.15 x 10-8 m2 for consolidated zone and 7.98 x 10-9 m2 to the other zone. The airflow pattern in the gob assigned with given permeabilities was expected to follow the experimental distribution (Brunner, 1985). The mean particle sizes for the semi-consolidated and consolidated zones were 0.02 and 0.006 m, respectively; smaller than those of the unconsolidated zone. Figure 2 shows the 3D view of permeability changes in the gob.

Jan 1, 2009

By Gene Mattila, Ron Boyd

Introduction Rock bolting is a general term that encompasses pinning, roof bolting, rib bolting, and cable bolting. Rock bolting contributes the highest nonproductive cost and is the most significant safety risk in mining today. Total installed bolt costs are now more than $20 per bolt. Rebolting or repair work doubles this. And bolting accidents account for about 30% of all fatal and nonfatal lost time accidents. It is only logical, then, for the mine operator to install the highest quality anchors possible. Every bolt on the market - tensioned, untensioned, full contact, or point anchor - has designed capabilities in axial, shear, and fluctuating ground conditions. Bolts will provide a designed support if material specifications and installation procedures fall within the manufacturer's recommendations. This also assumes that there are no unexpected geological or operational changes. Proper material specifications and installation procedures should guarantee the mine operator a quality anchor. To help assure this, the Mine Safety and Health Administration (MSHA), the US Bureau of Mines (USBM), and the AMC Manufacturers group developed stringent ASTM materials standards. These govern mechanical point anchor bolts and bolts using threaded and nonthreaded rebar. There are also minimum standards that suppliers must comply with in the manufacture of frictional bolts. However, most frictional bolt standards are based on the manufacturer's own specifications. It would be impossible to set uniform regulations on frictional bolts. Each bolt provides support in a slightly different manner and thereby requires different material needs. Quality Control In any case, material quality can generally be assured through tight order specifications and periodic quality control checks. The problem with installing quality anchors does not solely lie with the material. It includes all the events that occur from the time a bolt leaves a manufacturer to the time it is installed. The anchoring ability of any major type bolt can be affected by transportation, storage, mechanical performance of installation equipment, installation demands of the operator, proper drilling accessories, and local geology. Table 1 and Fig. 1 document the negative events that can affect each bolt and the degree to which it can be affected. Drilling a hole of oversized diameter, for example, is an operator controlled factor. It will adversely affect most bolts. An over-sized diameter reduces, or even eliminates, the anchorage capability of a mechanical shell. It will reduce the support capacity of a split set to nearly nothing. And it will cause a partial or incomplete grouting around the resin bolt. Frequently, due to bit wear, holes are drilled undersized in diameter and may not be properly cleaned out. In this case, it will be difficult for the operator to push the mechanical shell into the hole, thereby causing him to drill another hole. In cases where the bolt is obviously not providing anchorage, operators will probably drill another hole and place another bolt. Rock Bolt Regulations It is the situations where a bolt is improperly installed in the hole, but gives the appearance of a good bolt, that are frightening. In fact, some safety officials and engineers believe that only 50% to 60% of installed bolts are working. These bolts cause irregular ground load transfers in other bolts and overload them. This causes hazardous situations and overstressing. It is these bolts that fail prematurely 200 to 255 mm (8 to 10 in.) behind the face and possibly contribute to the high accident rate. The government and several manufacturers recognized the need to check the integrity of rock bolts. These programs are in different stages today. Tight torque regulations have been established for mechanical

Jan 4, 1985

By N. Sahin, K. Baris, A. Ozarslan

This paper aims to assess the current and future role of coal in energy strategy of Turkey and evaluates the compatibility of the policies to the EU energy policy and strategy. The Turkish government set targets to utilize more coal reserves of the country in next decades. However, the country is also in the process of becoming an EU Member State, hence, it is expected that the energy policies have to comply with the EU. Moreover, Turkey ratified Kyoto Protocol, thus the country should limit greenhouse gas emissions. Clean Coal, Carbon Capture and Storage (CCS) technologies are expected to become an important instrument for the mid-term reduction of CO2 emissions from coal-fired power plants. An alternative technique to the conventional CCS methods is CO2 sequestration by mineral carbonation. Turkey has significant resources of magnesium silicate minerals (olivine, serpentine) deposits which can be utilised in CO2 sequestration by mineral carbonation method.

Jan 1, 2011

By L. B. Hales

Expert control of grinding and flotation plants has been successfully used in the minerals industry since the 1970?s. The earliest of these systems were written in a hard-coded fashion in FORTRAN, BASIC or Pascal. Second generation systems were built using the first experimental expert system shells that were being developed in the artificial intelligence community. Later systems were deployed in expert system shells designed for real-time processing plants that also included the ability to model the process with neural network models and optimize setpoint selection through the use of genetic algorithms. Significant performance increases have been achieved using these systems, but in general, they suffer from the static nature of their rules and to a degree the process models also. Typical improvements quoted in the literature suggest throughput rate increases of 4 to 8 percent and recovery increases of 1 to 4 percentage points. In over 30 years of designing and installing expert systems we have never seen performance numbers less that these but have often seen numbers in excess of the high values. However, once the system has been installed and tuned up the performance usually does not improve much thereafter. Granted, the initial increases in throughput rate and recovery are significant but a question always lingers, that is, can the expert strategies be modified further to squeeze a little more improvement out of the system. The answer to this question is a resounding yes, however a complete understanding of what we do well as designers and operators of expert control systems is required as well as how we can improve upon our best practices is required to move us from the performance plateau we have been on for a number of years to a new higher level of performance.

Jan 1, 2009

The 112th Annual National Western Mining Conference (NWMC) hosted by the Colorado Mining Association (CMA) drew some political heavyweights, if by voice only, into the fray of two of the hottest topics that the mining industry is currently facing ? climate change and the new energy economy. A changing political landscape in Colorado, as well as in Washington D.C., has given rise to questions about the economic future of the United States and the mining industry?s role in that future. While there has been a recent push away from traditional energy sources to a ?new energy economy? that is rich in renewable energy sources such as wind and solar, the NWMC?s agenda emphasized the role that minerals will play in new energy technologies. The conference was also full of talks about the industry?s role in reducing the United States? reliance on foreign oil and other mineral commodities. To that end, an impressive lineup of speakers were invited to attend the three-day conference Feb. 8-11, in Denver, CO. Unfortunately, Mother Nature had other plans and three of the four headliners, including Denver?s mayor and Democratic candidate for governor of Colorado, John Hickenlooper, were unable to get to the conference because of a massive blizzard on the East Coast locked them in Washington, D.C. However, with modern technology, a full auditorium heard presentations from Hickenlooper, National Mining Association (NMA) president and chief executive officer Hal Quinn and Glenn Kelly, vice president, government affairs for NMA. Republican candidate for governor of Colorado, Scott McInnis, did speak to the crowd in person.

Jan 4, 2010

By Faruk Oksuz, Phillip L. Covell, Martin Doll

A series of difficult ground conditions causing high groundwater inflows and hazardous gas intrusion abruptly halted the construction of a 6.4-meter (21-foot) finished diameter and 1,890 meters (6,200 feet) long tunnel planned to serve as a 4-millioncubic meters (1-billion gallon) per day capacity river outfall for treated wastewater. Some 808-meters (2,650-feet) section of the constructed tunnel, TBM, and all other tunneling equipment were completely flooded with hydrogen sulfide laden artesian groundwater. The subsequent pumping and tunnel recovery efforts were unsuccessful and the tunnel was abandoned. In the end, the construction contract was terminated and options to redesign and reconstruct the project were re-evaluated to maximize the use of already constructed tunnel facilities such as the construction access and inter-mediate shafts and six outfall diffuser shafts in the river.

Jan 1, 2007

By M. L. Free, L. Anthian

Electrochemical, chemical, and image analysis techniques were utilized together with statistical analysis methods to determine the effect of various anode impurities on copper electrorefining performance. Results indicate that arsenic is a beneficial anode impurity. Anode lead and sulfur are shown to decrease electrorefining performance. Specific emphasis is given to the influence of individual elements on current efficiency, slime particle generation, and cell voltage parameters.

Jan 1, 2000

By H. C. Seeber

When considering all the aspects of a viable mine plan, it is important to estimate the accurate cost of mine ventilation, which varies as a function of airway dimensions, fan selection and ventilation network. Optimization of the overall elements of ventilation design can lead to very significant cost savings when capital and operating cost savings are considered in combination. This paper will describe a methodology, whereby ventilation costs can be minimized by optimizing the ventilation aspects of mine design and fan selection, by providing a brief description on ventilation theory and the factors which affect ventilation operating and capital costs.

Jan 1, 2002

Many mining operations generate seismic signals. They come from blasting and from underground mine failures such as rockbursts, longwall first caves, coal bumps and pillar collapses. Traditionally, seismic motions have been of concern on a local scale because they could create damage to structures at the mine site or at neighboring locations. Recently, there has been much interest in the Comprehensive Nu-clear-Test-Ban Treaty (CTBT). It prohibits any kind of nuclear explosion, anywhere. So distinguishing signals of mining operations from a possible nuclear explosion becomes important. An international monitoring system (IMS) is being implemented to verify the Treaty when it is ratified. It includes a large number of seismic stations capable of recording signals at regional (200 to 2,000 km or 124 to 1,240miles) and teleseismic (more than 2,000 km or 1,240 miles) distances.

Jan 1, 2000