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By V Dolmage

In 1957 and 1958, ten topographically suitable damsites were investigated along the Peace river between Finlay Forks and Hudson Hope, B.C. Of these, three were eliminated after a reconnaissance survey indicated that the geology was unsuitable for the 650-ft.-high dam contemplated. ?After extensive drilling, five other sites were deemed unfeasible because of geological factors and, in 1959, the two remaining sites were designated as being suitable, one for the high dam and the other for a low dam downstream. The sites are located in the Peace River canyon, on the west and the south? side of Portage mountain, west of Hudson Hope. The site for the large dam, the Portage Mountain dam, has since been investigated by surface mapping, thousands of feet of diamond drilling, a seismic survey, laboratory rock tests and on-site grouting tests in order to establish the possible effect on the dam of all facets of the foundation rocks. Three diversion tunnels, 50 ft. in diameter and 2,200 feet in length, have been driven through massive sandstone and shale, and an intake cut, 500 ft. in length, 200 feet in width and 200 feet in depth, has also been excavated in rock.

Jan 1, 1963

By K. Chomyn, R. Veenstra, P. Zangeneh, Hatch Ltd., C. Van der Woude, H. Ghorbani

"The furnace is the heart of smelting operations, and therefore the performance, reliability, and campaign life of the furnace are critical considerations. Furnace refractory is exposed to high temperatures and other harsh operating conditions causing brick erosion, crushing, and cracking, typically limiting the furnace campaign life and requiring costly furnace shut-downs for refractory re-lining. Despite extensive study, the in-service thermal / mechanical stresses in these complex refractory systems are not generally well understood. Hatch has developed a thermal-mechanical finite element analysis (FEA) based furnace refractory modeling approach that allows for the quantification of refractory movements, load distribution, and thermal / mechanical stresses.Due to the complexity of the refractory system, a range of models are used to identify and assess damage mechanisms and design improvements, including: 2D models, to capture expansion and movement of individual bricks, changes to load paths, and thermal stresses; and 3D models, to include the effect of differential expansion between bricks in each course, and assess overall load distributions. The 2D and 3D analyses include the effect of nonlinear temperature-dependent properties, thermal expansion, movement of individual bricks with frictional contacts (i.e., gaps can open, bricks can slide), and brick tensile or compressive failure. Both as-built and damaged conditions can be investigated (e.g., after bath zone brick erosion), for circular and rectangular furnaces. 3D models also include papering and brick joint & interface action.Analyses using these models have shown the importance of thermal stresses, papering, and geometry / arrangement of individual bricks. Results have been used to identify and assess damage mechanisms, showing good alignment with furnace autopsy observations. By using this FEA-based approach to quantify the refractory design, the risk of damage can be significantly reduced, low-cost design changes can be identified, and the furnace campaign life can be extended.INTRODUCTIONThe furnace is the heart of smelting operations and therefore the performance, reliability, and campaign life of the furnace are critical to the success of the smelting operation. Furnaces are inherently exposed to high temperatures, often in excess of 1600°C (Holappa, 2004), and furnace operating power has progressively increased over time (Voermann, Gerritsen, Candy, Stober, & Matyas, 2004). The refractory system must withstand high temperature, high heat flux, and other harsh operating conditions. These conditions are particularly significant for the furnace hearth, and can lead to damage including brick erosion due to molten material flows, crushing due to thermal expansion loads, and cracking due to severe thermal gradients within the bricks. Refractory damage typically limits the furnace campaign life, requiring costly furnace shut-downs for refractory re-lining (repair or replacement). In some cases the refractory may lose integrity causing damage to nearby equipment and harm to personnel."

Jan 1, 2016

By T. I. . I. Sibbald, D. E. Jiricka

The Goldfields area on the north shore of Lake Athabasca , was active in the thirties and early fortie . In 1934, Tom Box discovered gold mineralization in quartz veins in pyritized 'granite' at the ite of what wa to become the Box Mine of the Consolidated Mining and Smelting Company of Canada Ltd . In the rush of activity that followed numerou other gold occurrences were found. the most significant of the e being the Athona and Frontier prospects which underwent underground development bur did not reach production. Probable reserve identified to date in the Box and Athona deposits amount to some 3.5 million ton grading an average 0.ITT oz gold/ton . Some 10 to 15 million tons of lower grade, po ible reserve are also reported at the Box. From 1939 to 1942, the mine produced, from underground , 64,066 ounces of gold from ore grading 0.0452 oz gold/ton at a cost of between $1.5 and $2.0 a ton. The Goldfield area is underlain by upracrustal rock intruded by mafic bodies and invaded by ·granite'. The upracru tal . thought to be of Lower Proterozoic age are characterized by a helf-type ucce ion metamorpho ed under lowermiddle amphibolire facie conditions into quartzites, quartz-mica chi ts. dolomitic marbles, diop idic calc-silicates and cordierite-anthophyllite-biotite schi t . Amphibolites of presumed igneou origin form concordant to subconcordant units up to a few ten of metre thick in the upracrustals and occur as le regular masses in the 'g ranite·. Compositionally they vary from gabbroic to granodioritic. with the more mafic varieties being most common. Granitic rocks exhibiting a variety of textural , structural and compositional characteristics underlie most of the area. Four main units are recognized: foliated (older) 'granite', 'leucogranite', feldspar augen 'granite· and pyritic mine 'granite', however, in geological mapping the e are somewhat arbitrarily differentiated due to transitional relationships and admixture of types. Although there is no unambiguous age data available for the e 'granite ' adjacent northshore 'granite' of geologically similar age yield Rb-Sr isochron dates around 2000 Ma. The area is characterized by two major deformation events of which the earliest gave rise to the dominant regional foliation and to rare minor and outcrop scale folds. This foliation which in general parallels compositional layering is folded into the major Goldfields syncline, whose axis plunges weakly to moderately outh. Related open minor folds have subvertical axial surfaces which are locally paralleled by a weakly defined schistosity. Most rock display a pronounced rodding fabric, coaxial with second fold axes and apparently the product of this deformation event. However, a note of caution is introduced into this interpretation, as the rodding fabric is also parallel to first fold axe .

Jan 1, 1986

Electric Vehicle Controllers Ltd. ?EVC started in 1981 ?Underground electric locomotive drive systems ?Developed tandem load sharing system and radio ELOS controls ?Locomotive autonomy systems ?Entered space technology development in 1999 with NORCAT to develop Mars drill for CSA ?Commercial distributor for drilling systems, lunar physical simulants, electric mobility systems and lunar mining technologies

May 1, 2009

In the last 30 years numerous mining plants have been developed in North America but only a handful of these have required freezing the area where the shaft was sunk. The most recent mine shaft to be sunk using this technique is located on an island off the coast of Louisiana. The project required the input and skills of numerous people in several companies, including the owner, engineers and contractors. The Cote Blanche installation is one of several mines in the area which extract salt from a dome which extends under the Gulf of Mexico. The geology of this area is, however, markedly different from others in the region because there is approximately 460 ft of water logged soil above the salt dome. Because it is not possible to sink a shaft in ground that is this soft, the overburden where the shaft is located must be frozen before excavation can take place. The Cote Blanche mine is a producing facility which was built in 1958 by the Carey Salt Company. The original facility included a 14 ft diameter production shaft and a 7 ft-8 in diameter service/ventilation/escape shaft. This system was installed by The Cementation Company. In 1996 the owner decided to improve both the mine ventilation and production hoisting system by installing a third shaft. This shaft would be designed to extract 650 tons per hour from a depth of 1600 ft and provide 500,000 cfm of ventilation to the mine at a maxium design pressure of 4 in WG. These parameters resulted in a requirement for a 16 ft finished inside diameter concrete lined shaft and a headframe designed as a pressure bulkhead. Because the mine would operate from one level only, the most efficient hoisting system for the application was determined to be a ground mounted koepe running on rope guides. With these criteria fixed Cote Blanche embarked on a program which would see construction start in June 1998 with the first salt produced in February 2001 at a cost of $US18M. The major players in the design and construction were G.L. Tiley & Associates Ltd., BLM Bharti Engineering, Freezewall Inc., Phillips Mining, Geotechnical and Grouting Inc., Don Hart Consulting and Dynatec International.

May 1, 2002

By J. Lu, M. Mohammadi, C. Pelletier, B. Wassink

Chemical reduction of selenate by chromous ions is proposed as a novel method to remove selenate from sulfate-bearing solutions Chromous ions were generated by electrochemical reduction of chromic ions using an electrochemical cell. The investigation was conducted in sulfate medium and under various concentrations of chromous ions (0.0062 to 0.05M), pH 1 to 2.2 and temperature 10 to 60°C. Selenate ions were rapidly reduced by chromous ions to hydrogen selenide. The overall reaction at a chromous/selenate molar ratio of 8 or higher can be expressed by the following reaction. At an initial chromous/selenate molar ratio below 8, the products are hydrogen selenide and elemental selenium. At a lower chromous/selenate molar ratio and pH, more elemental selenium and less hydrogen selenide are produced. The reduction kinetics is controlled by both chromous ion and hydrogen ions concentrations. The reaction rate is first order with respect to selenate, hydrogen ion and chromous ion concentrations. It was found that the selenate reduction rate with chromous is independent of the concentration of sulfate present in the solution. The general rate equation was proposed and the rate constant was calculated as k= 0.38 M-2s-1. An activation energy of 58.7 kJ/mol was obtained in the temperature range from 10 to 50°C. Nearly 100% of selenate can be reduced to H2Se from solution containing 17 ppm selenate in less than 30 minutes at pH 2.2 and 60°C. Various strategies have been tested to remove the produced H2Se from solution. INTRODUCTION Selenium in nature is mostly associated with metal sulfide minerals such as copper, iron, zinc, and lead sulfides (Zingaro & Cooper, 1974). The extraction of metals from ores has the potential for selenium contamination of ground and surface water which can pose an environmental and health concern. Therefore, selenium contamination in the receiving environment is a key issue for many industries (specifically copper industry) and the treatment of wastewater contaminated with selenium is a growing challenge. The effective removal of selenium from contaminated water is very important to support sustainability in the copper industry and the other industries with selenium pollution issues.

Jan 1, 2019

By A. Taylor

This paper considers mineral processing in zero gravity dry environments typical of asteroids or comets. The electro-dynamic separation technique using pneumatic flow systems is explored. Pneumatic delivery of solids is used because of extreme temperature changes and the possible scarcity of water. The present earth-based process and how it can be modified for these conditions is examined.

May 1, 2001

By Leopoldo Gutierrez, Sergio Castro, Lina M. Uribe, Janusz S. Laskowski

"The effect of sulfite and metabisulfite ions on the amyl xanthate flotation of copper-activated pyrite, in distilled water and seawater, has been studied by micro-flotation tests and redox potential measurements. The activation of pyrite was carried out varying copper ions concentration. It was found that in the presence of copper ions in solution, sulfite and metabisulfite are able to depress copper-activated pyrite at neutral or slightly acid pH. After the addition of sulfite or metabisulfite ions the resulting pH is more acidic with metabisulfite ions. A significant increase of the redox potential of the solution was observed as copper ions concentration increased. Then, the initial reducing conditions brought about by sulfite and metabisulfite ions, change abruptly to oxidizing conditions in the presence of copper ions. Under these oxidizing conditions, the metabisulfite showed to be an effective depressant for pyrite mainly in seawater. However, sulfite was a pyrite depressant in both distilled water and seawater. Based on these results, the pyrite depression by sulfoxy-depressants during the flotation of Cu/Mo ores in distilled water and seawater is discussed.INTRODUCTIONDepression of pyrite by sulfoxy-depressants (sulfite and metabisulfite) in the presence of copper ions is an important issue. This is particularly true when pyrite depression in the flotation of Cu/Mo sulfide ores is carried out in seawater. The use of lime in seawater must be avoided because the formation of magnesium hydroxy-complexes and precipitating magnesium hydroxide strongly depresses molybdenite flotation (Castro et al., 2014). In order to solve this problem, one large plant in Chile uses sodium metabisulfite (MBS) instead of lime to depress pyrite in seawater. In plant practice, this also improves molybdenite flotation since with the use of MBS the pyrite depression is carried out at pH 6-7, instead the pH 11-12 usually employed with lime.From a general point of view, pyrite exhibits some natural floatability over acidic pH ranges and may be partially floated without thiol collectors (a coating of sulfur/polysulfides has been claimed). Therefore, pyrite flotation with xanthates is usually high in the acidic pH region, but decreases greatly at alkaline pH (Leppinen, 1990; Fornasiero and Ralston, 1992; Zhang et al., 1997)."

Jan 1, 2016

By Karl J. C. Harries

"6.1. INTRODUCTIONThis chapter comments on some of the more common types of agreements related to property examination and acquisition, other than the farm-in/joint venture agreement or common undertaking, which is dealt with in the next two chapters. Many of the comments made in Part IV of Chapter 4 apply equally to any such agreement. In addition, many of the comments in Chapter 8 and the examples in Volume II are most certainly relevant to many other types of agreement.6.2. RIGHTS OF FIRST REFUSAL1A right of first refusal is a right that obligates an owner to offer his property, or an interest therein, to the holder of the first right if the owner intends to sell or otherwise dispose of it. It may be triggered either by the owner wishing to offer his property for sale or by receipt of a bona fide offer to purchase his interest from an “outside” third party. The mechanics and procedures to implement a first right should be set out clearly in the agreement, and their basics are the same, whether the first right is a stand-alone document or is part of an acquisition agreement.2Rights of first refusal are common in agreements to permit some control over the disposition of an interest and there they can be a valuable provision. However, some explorationists feel that they are also a good stand-alone agreement to have while a property is being evaluated to decide whether or not it is worth optioning. The reasoning is that such an arrangement will tie up the property to allow a quick look and, if the property looks interesting, the balance of the term of the first right will give an ample exclusive period of time to negotiate with the owner and settle an acquisition agreement. In addition, the explorationist is protected by the first right if the owner decides to negotiate a deal with another party."

Jan 1, 2003

By H. Sakurai, M. Kato, F. Tanaka

The Mitsubishi Continuous Copper Smelting/Converting Process employs an electric furnace (CL- furnace) for settling the slag/matte mixture formed in the smelting furnace. The conditions in the CL- furnace change its settling efficiency, resulting in high variability of copper losses in slag. Modifications of the operation and maintenance procedures for the CL-furnace have significantly and sustainably reduced the copper losses in Gresik smelter to below 0.7%. A series of flow patterns evaluation in the CL-furnace demonstrated how the chemical reactions and settling conditions of the CL-furnace has been improved by the new operating procedures. Settling mechanisms for the matte droplets in the CL-furnace is also discussed using image analysis of solidified slag. INTRODUCTION The Mitsubishi Continuous Copper Smelting/Converting Process employs an electric furnace (CL- furnace) for settling the slag/matte mixture formed in the smelting furnace. Since the top-blowing in the S- furnace agitates the molten bath strongly, matte droplets are dispersed over the slag phase when melts flow out of the S-furnace. Thus, the CL-furnace has an important role to settle and recover matte droplets efficiently. Since many papers have ascribed major part of copper losses to matte droplets suspended in slag (Tanaka 2007; Genevski et al., 2008; Starodub et al., 2016), flow patterns of slag in the CL-furnace can affect greatly on the profitability of smelters which have adopted the Mitsubishi process. PROCESS DESCRIPTION Figure 1 illustrates the Mitsubishi process schematically. The Mitsubishi process is the only system for continuous production of blister copper from copper concentrates and consists of 3 furnaces connected by enclosed launders those ensure the minimum emission of fugitives. In a smelting furnace (S-furnace) copper concentrates are fed with flux, oxygen-enriched air to form a mixture of slag and matte which cascades down to a slag-cleaning furnace (CL-furnace). The CL-furnace settle matte droplets in the mixture on the crucible from which matte is siphoned to a converting furnace (C-furnace), while discharging the cleaned slag (CL-slag). The C-furnace converts matte to blister copper by the top-blowing of oxygen- enriched air.

Jan 1, 2019

By N. Schreithofer, M. Aspiala, R. Serna

The novel attachment timer apparatus presented in this paper serves as a diagnostic tool that allows the detection of changes in bubble-particle attachment probability and therefore floatability caused by alterations in the chemical environment and particle properties. The apparatus consists of six identical capillaries where bubbles with defined size are produced simultaneously in a measurement chamber. The bubbles at the capillary tips are placed in contact with the submerged particle bed for specific time periods, controlled with the help of automatic actuators. The advantage of the instrument is that hundreds of bubble-particle contacts can be measured automatically within a short time period. Microscopic pictures of each measured bubble are taken while recording the movement of the bubble before, during and after contact with the solid particles. The recorded pictures can be used to determine the actual bubble size and its corresponding deviation, and to detect the attachment of particles. The attached particles are collected in a detachable chamber for subsequent characterization. Furthermore, the device is portable and can be taken to the mineral processing plants for quick evaluation of particle-bubble attachment efficiency with particles and process water sampled directly from real processes.

Jan 1, 2017

By W. L. Brown

"The properties of Base Metal Mining Corporation are in Yoho Park about 2 1/2 miles northeast of Field; British Columbia; on the main line of the Canadian Pacific railway. There are actually two mines; the Monarch on mount Stephen and the Kicking Horse on mount Field; separated by the steep valley of Kicking Horse river. The outcrops of the orebodies range from 4; 900 feet to 5; 300 feet above sea-level. The concentrator is situated below the Monarch mine and is connected to the main portal by an aerial tramway. The ore from the Kicking Horse mine is conveyed by aerial tramway to a storage bin on the main highway and is then trucked to the concentrator.During the construction of the Canadian Pacific railway through Kicking Horse valley, galena float was found at the base of mount Stephen. The East Monarch outcrop was staked in 1894. The outcrop of the West Monarch ore body on the cliff face of mount Stephen was not discovered until 1916. The Kicking Horse orebodies on mount Field were discovered a few years later. In 1912, a 75-ton concentrator was erected, which operated at intervals until 1924. During this period approximately 42,000 tons were mined and milled, largely from the East Monarch orebody. In 1928, the Base Metal Mining Corporation, Limited, was formed to take over the Monarch and Kicking Horse mines. A 300-ton differential flotation concentrator was erected and production started in the autumn of 1929. In a very short time, however, the low price of metals forced the mill to shut down until August 1933. From 1929 to 1935, when the mill again ceased operation, a total of 256,500 dry tons of ore, averaging 1.59 oz. silver per ton, 9.7 per cent lead, and 12.6 per cent zinc, was mined and milled, largely from the West Monarch orebody. After a period of further underground development, the concentrator reopened in 1940 and is still in operation. From 1940 to the end of 1944, 354,363 tons of ore was milled, averaging 5.32 per cent lead and 10.47 per cent zinc."

Jan 1, 1954

By J. E. Hawley

"The property of Francoeur Gold Mines, Limited, lies directly west of the Armfield mine in Beauchastel township, Abitibi district, and 12 miles west and slightly south of Noranda. The gold deposits are of particular interest because of their replacement character and close association with fault zones. The property was originally staked in 1923 following the discovery oi a mineralized zone near the eastern boundary. This became No. 1 orebody. Later discoveries in 1936 of No. 2 and No. 3 orebodies led to underground development and actual mining in July, 1938. A fourth orebody, No. 8, was found in 1939.The following is the production record up to December 31st, 1947:Beginning in September, 1944, ore was shipped to Noranda as siliceous flux, but late in 1946 milling of low-silica ore from No. 1 orebody was undertaken until operations were temporarily suspended in 1947.The four main ore deposits, Nos. 1, 2, 3, and 8, consist of siliceous ferrocarbonate and albite replacement bodies mineralized with fine pyrite and specularite in schistose fault zones in Keewatin volcanics. Barren to low grade quartz veins though present in some are uncommon, but small stringers of quartz are usually present wherever here and there visible gold is found. Nos. 1, 2, and 3 deposits lie in a major fault zone that extends completely across the property from west to east (over 7,000 feet) and dips N.40° to 50°, and are separated by intervals of 1,800 to 2,000 feet respectively. The fourth lies in a smaller fault zone striking east and dipping on the average 65°S.No. 1 orebody has been explored by mining to a depth of 500 feet, No. 2 to 400 feet, and No. 3 to slightly less than 300 feet . The greater part of No. 8 lies above a depth of 200 feet, but a raise has recently .been driven through from the west end of the 5th level of No. 1 to the 2nd level on No. 8 zone. Most of the ore of No. 8 zone above the 2nd level and of Nos. 2 and 3 above the 3rd levels has been removed."

Jan 1, 1954

By Michael Samis

Conventional project valuation techniques, such as discounted cash flow (DCF) method, are more frequently being criticised for the manner in which they consider the interacting effects of uncertainty and project structure (including management flexibility) on project value. Modern asset pricing (MAP; often called ?real options valuation? when applied to the analysis of flexibility) is a more general valuation framework that corrects many of the shortcomings of common valuation methods. Previous mining applications of MAP have tended to treat the ore deposit as a homogenous entity. While useful for demonstration purposes, this is inadequate for practical application in industry. This paper presents a MAP valuation example that allows for the heterogeneous nature of an ore deposit. It shows that deposit structure may increase the complexity of the mine planner?s decision so that results of a standard DCF analysis, or a simple MAP analysis that ignores the details of deposit structure, are misleading. In the example, a management is considering the development of an adjacent low-grade zone while mining operations in a high-grade zone are ongoing. If the low-grade zone is developed before high-grade reserves are exhausted, the proposed development will include both low-grade zone development and a capacity expansion to handle additional low-grade reserve production. If development of these reserves is delayed until the high-grade reserves are depleted, then the low-grade reserves can be developed without a capacity expansion. A typical DCF analysis would reduce the management?s decision problem to three alternatives: 1) continue mining only in the highgrade reserves, 2) immediately develop the low-grade reserves and expand capacity, and 3) delay low-grade zone development until the high-grade reserves are exhausted. The model presented here allows the management to reconsider the development and capacity expansion decision on a discrete half-year basis. It delineates price regions, for given levels of high-grade zone reserves, in which it is optimal to develop the low-grade zone, defer low-grade zone development and abandon the mine irrevocably. The model presented here extends the ability of mine planners to describe actual project structure while using the MAP valuation framework that is already noted for its capacity to combine project uncertainty and risk insights with a dynamic project description.

May 1, 2001

By R. A. (Dick) McIvor

"It is important to ensure that the basic mine hoist plant design fundamentals are incorporated into any new mine hoist plant project before detailed design engineering is commenced. Some of the basic design fundamentals to be considered are: hoist rope fleet and flight angles for good hoist rope coiling, hoist rope harmonic frequency study to ensure that the hoist rope does not run at full speed at a rope resonance frequency, head frame and shaft clearance diagrams to establish the optimum positions for the loading and dumping of the shaft conveyances, access in the head frame and shaft collar for hoist rope and shaft conveyance installation and removal. Failure to consider the hoist plant fundamentals before detailed engineering commences will result in extra costs for hoist plant redesign or a significant increase in operating and maintenance costs for the mine due to increased hoist plant downtime to perform tasks and a lower mine operating efficiency. INTRODUCTION The mine hoisting plant is the most critical single line component in the operation of a shaft sinking operation or a permanent installation. Bad decisions in the design and selection of the hoist plant equipment can significantly add to the operating costs and up-time in the operation of the mine. Mine hoisting plant equipment installed costs represent approximately 5 to 8 percent of the capital cost of a new mine, yet price sometimes becomes the deciding factor for procuring the hoist plant components, sometimes selecting used equipment that may compromise the design and maintainability of the mine hoist plant. Mine downtime can cost upwards of $100,000 per hour and it will not take a very long time to eat up all the so-called savings on hoist plant equipment purchases. This paper will provide guidelines for the design, selection, installation and maintenance of mine hoist plant equipment with the objective to having the right mine hoist plant with the highest operational availability and best maintainability. PRELIMINARY DESIGN INFORMATION Before any engineering commences, the following information needs to be obtained to allow the preliminary sizing of the hoist plant equipment: • Shaft depth: The shaft depth will influence the hoist rope selection based on the material and personnel loads anticipated for the mine design. The shaft depth will also influence the maximum speed of the shaft conveyances. • Annual production: The desired annual production will influence the required payload sizing with allowances for maintenance and inspections of the hoist plant equipment."

Jan 1, 2019

By E. Brzychczy

Artificial intelligence techniques are ever more frequently implemented in the modern world. One of their numerous uses is supporting enterprise management. Intelligent techniques and intelligent information technology (IT) systems can be implemented in mining-related issues as well. The author of this article concentrates on supporting the process of planning mining operations by intelligent systems and techniques. The author also underlines the importance of knowledge accumulation in a mining enterprise (especially tacit knowledge) as well as its efficient storing and using. Additionally, the author provides a short description of characteristics of intelligent IT systems and indicates their capabilities in planning mining operations. She presents two solutions, namely: an advisory system which supports the process of planning preliminary works and exploitation works in hard coal mines, and a method of modeling and optimizing mining operations. Finally, the author indicates a few possible future development trends of intelligent systems - the intelligent systems of decision support.

Aug 1, 2013

By Jan Butra

Longstanding mining of copper deposits in the Legnica-Glogow Mining Basin, Poland has accumulated broad experience concerning the impact of geological and mining conditions on seismic and rockburst hazard. Methods of rockburst prevention for deep copper mines have been developed based on gathered data. Methods of rockburst hazard evaluation and rockburst control constitute a set of rockburst prevention measures. Due to their nature and location within the production process, prevention methods can be passive or active. Active methods of rockburst prevention consist of the liquidation of stresses in the areas where they are concentrated through controlled generation of rock-mass tremors and rockburst by blasting works (group extraction blasting, release blasting in the rock mass and pillars able to accumulate the stress, as well as in the roof and floor). Group blasting works as a basic method are very effective in copper mine conditions. Provocation of dynamic events makes it possible to control the timing of their occurrence. In Polish copper mines, all blasting works are made during the absence of miners in the mining fields. Daily distribution of both the number of recorded strong seismic events and volume of energy released, confirm the advisability of using this kind of rockburst hazard prevention. The rockburst hazard prevention methods developed over 40 years of copper deposit extraction are economically reasonable and offer sufficient effectiveness under high rockburst hazard conditions.

Aug 1, 2013

By M. Nzotta, L. Teng, S. Seetharaman, A. Chychko

The objective of the present work is to optimize the EAF practice towards Mo alloying with maximum retention of the alloying element in molten steel. During Mo additions, there were significant losses to the dust (ca 7 %) and to the slag phase (ca 2 %), which had serious economic and environmental impacts. The present work was aimed at designing a suitable Mo precursor, an alternative to ferromolybdenum, as the Mo source and a process sequence for optimal retention of Mo in the steel. The design of the Mo source should be in such a way that low cost raw materials such as MoO3 concentrate have to be used, stable during the addition so that the vapour phase losses as MoO3 are kept minimum and the dissolution of the precursor in molten steel is easy without loss to the slag phase. After thermodynamic considerations, three precursor materials were chosen, viz. CaMoO4, MgMoO4 and Fe2MoO4. Synthetic mixtures of the component oxides in each case as well as pure MoO3 were investigated by TGA and high temperature XRD in order to find out the temperature range in which the compounds were formed. Experiments were also conducted separately where the evolved gases were analyzed. All the three precursors were formed easily and the loss of MoO3 to the vapour phase was found to be extremely low. Addition of the oxide mixtures to steel melts were conducted in three different scales, viz. 16 g, 500 g and 3 tons (induction furnacle). Initial experiments showed that MgMoO4 was not suitable since the Mo yield in the steel melt was not very high. Fe2MoO4 was found to perform best, the Mo yield reaching 98 %. Further trials were conducted in the 65 ton Electric Arc Furnace (EAF) at Uddeholms AB in Hagfors. Mixtures of mill scale, MoO3 and carbon in suitable proportions were added in a pre-determined addition sequence. It was found that, in the actual practice, the Mo yield reached 99 %, equaling in performance the addition of ferromolybdenum. The process offers an effective use of cheaper raw materials and very high Mo yield avoiding a pre-treatment step like the production of ferromolybdenum. This is both cost effective and environmentally advantageous. The process has been patented and is being currently implemented at Uddeholms AB, Hagfors, Sweden.

Jan 1, 2011

By Charles S. Ney

"THE WRITER'S acquaintance with Kootenay King mine is limited to a brief geological excursion in the vicinity. This paper has been compiled largely from maps of J. Sullivan, mine geologist in 1952, kindly made available to the writer by Transcontinental Resources; a 1942 private report by W. L. Brown; and personal communications with J. Sullivan. Remarks on regional geology, and the comparison with Sullivan mine, are specifically the writer's.The Kootenay King mine is situated in the westernmost range of the southern Canadian Rockies at an elevation of 7,300 feet. From the Sullivan Mine at Kimberley, B. C., it lies 19 miles due east. Six miles of this distance are represented by the pronounced topographical and structural gap of the Rocky Mountain Trench. Mineral claims were locat-ed in the area prior to 1898. The deposit was first seriously explored in the late 1920s, and a small orebody was then outlined and developed. Mining and milling were undertaken in 1951.52. The orebody, amounting to about 14,000 tons, was mined out, and no significant new ore was revealed.The orebody is a replacement of dolomitic argillite by fine-grained galena, sphalerite, and pyrite. The sulphides are finely laminated in the direction of bedding, following out minor folds faithfully. The grade of ore estimated from diamond drill holes (W. L. Brown, 1942) was 11 ·I per cent zinc, 11 ·I per cent lead, and 3·8 ounces per ton of silver. The mineralized zone extends about 250 feet horizontally, 160 feet vertically, with a width of six to twenty feet.General GeologyThe country rocks are sediments of the Aldridge formation of late Precambrian age. For several thousand feet stratigraphically above and below the mine, the formation consists mainly of thinly laminated argillite and siltstone. Several horizons contain appreciable carbonate. About 2,800 feet stratigraphically below the mine, predominantly dolomitic beds mark the top of the underlying Fort Steele formation. The medium-bedded fine quartzites, which characterize about 10,000 feet of the Middle Aldridge in the Purcell range, have a restricted development amounting to only 2,000 feet in the Rocky Mountains. This member is about 2,200 feet above the mine horizon."

Jan 1, 1949

Fiabilité Totale d?Usine (FTU) Plan de la présentation ?Théorie de l?entretien ?Analyse comparative juin 2005 ?Session de travail 2005-2006 ?Suiviet Résultats(IndicateursdePerformance)

Feb 1, 2008