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By Kanaya Koichiro

Doyashiki Concentrator in Hanaoka mine of Dowa Mining Co., Ltd. started the operation in 1943, treating the ore mainly from Doyashiki ore deposits. After the cessation of mining of Doyashiki ore body in 1973, Doyashiki concentrator began to treat the ore from Fukazawa mine instead in the same year, and furthermore the ore from Ezuri mine In 1979. Currently Doyashiki concentrator treats 30,000 tons of ore per month from both mines. The ore treated is complex sulfide ore, called "Kuroko" (black ore), and the principal minerals contained in the ore are chalcopyrite, galena, sphalerite, pyrite, and barite, accompanied by bornite, tetrahedrite, tennantite, covellite, stromeyerite, argentite, polybasite, pearcite, and electrum. The Principal gangue minerals are quartz, gypsum and clay. The grade of the feed to the concentrator averages 0.9 p/t Au, 110 g/t Ag, 1.1 %Cu, 2.8 %Pb, 9.2 %Zn, 5.0 % Pyrite-S and 13% BaS04. The feed is, after washing and two-stage crushing, ground to 88% minus 200 mesh in a rod mill and ball mills with cyclone classifiers, and five concentrates are recovered by flotation. Primarily copper-lead bulk concentrate is recovered by sulfurous acid-lime flotation method, and then from the bulk concentrate copper concentrate is floated by hot pulp flotation method, sinks being lead concentrate. Zinc concentrate is recovered from the tailing of copper-lead bulk flotation, pyrite concentrate from zinc flotation tailing and barite concentrate from pyrite flotation tailing. The final tailing is sent to the dam.

Jan 1, 1982

"With up to 2,300 m (7,545 ft) of rock on top of it, the Gotthard Base Tunnel, which is scheduled to go into regular operation in December 2016 in the Swiss Alps, is not only the world’s longest, but also the deepest railway tunnel ever built. Accordingly, the temperatures in the tunnel are very high, with the surrounding rock expected to reach 45° C (113° F) once the tunnel is fully operational. Since unrestricted rail operations are only permitted in temperatures of 40° C (104° F) or less, the heat has to be controlled with suitable equipment. Large fans developed and built especially for this application by TLT-Turbo GmbH provide fresh air for regular operations as well as during maintenance and repair work. Separate fans ensure that the emergency exits are properly ventilated in case of fire. In addition to coming up with a technical solution that could handle the special challenges of this project, the engineers and planners had to develop detailed preliminary plans as well as a sophisticated logistics and installation concept.The Gotthard Base Tunnel consists of two 57-km (35-mile) long tubes, one in each direction. Including all connecting and access tunnels, the entire underground system measures 152 km (95 miles). At Faido and Sedrun, at approximately 16 and 36 km (9 and 23 miles) respectively, two multifunction stations were built that cut each of the two tubes into three sections of roughly equal length. Here, trains can make emergency stops and change tracks, if necessary. This is also where eight large axial fans made by TLT were installed that exchange the air, i.e. they extract a certain amount of heated air and inject the same amount of cool fresh air a few hundred meters further down the train tracks. In addition to the four exhaust and four supply fans, 24 so-called jet fans were installed along with associated components like silencers, shut-off dampers, drives and lubrication systems from TLT.In addition to exchanging the air, these fans are used to enable maintenance and repair operations in the tunnel. Current plans call for blocking each tunnel tube for one night per week in order to perform preventive maintenance work, thus enhancing the tunnel’s overall functionality and availability. During these hours, many people will be in the tunnel who require fresh air. “One of the major criteria for winning this bid was the fans’ ability to quickly switch between a wide-range of modes. TLT fans are among the best in the world in this regard,” said Andreas Kuhn, TLT’s project manager for the Gotthard Base Tunnel, about the selection of his company’s fans for this project."

Dec 1, 2016

The surface of chalcopyrite was studied in situ in aqueous solution by an atomic force microscopy (AFM). The AFM images showed that some absorbate in patches was detected on the mineral surface, after the chalcopyrite sample was soaked in 5x10-4M potassium ethyl xanthate (KEX) solution at pH 11 for 10 minutes. These patches increased in size with time elapsing. At the same time, as shown by the obtained AFM force curves, the adhesion between the AFM tip and the substrate increased greatly due to the adsorption. Rinsing with ethanol removed the absorbate from the chalcopyrite surface, as well as the big adhesion observed in KEX solution.

Jan 1, 2010

By G. Le Bel

Discussion by G. Le Bel The paper proposes three methods to determine optimum lifetime of a mining project 1. The conclusion is that the three methods arrive at the same optimum production rate (by coincidence). It is also stated that: "The stochastic and option pricing models, in addition to providing a more realistic optimum mine life than the traditional NPV (net present value) calculation, showed that the example project has the potential for substantially higher return. By explicitly considering a project's potential risks, the stochastic or option pricing techniques can more accurately describe its probable investment performance." This discussion is intended to demonstrate that: •It is no coincidence that the three methods arrive at the same optimum production rate. • The stochastic and option pricing models do not provide a more realistic optimum mine life than the traditional NPV calculation. •The potential for a substantially higher return is misrepresented. First, a review of investment decision theory in mining is needed. Net present value is the preferred performance measure because of the flexibility offered when comparing various alternatives. However, it presents the following disadvantages: • The discount rate selection is not straightforward. • The NPV obtained represents a deterministic characterization of the variables included in the cash flow calculations, i.e. a single-point estimate. Discount rate selection The opportunity cost of capital represents the rate of return that would be obtained if the capital was instead invested in other opportunities considered to exist. Such opportunities must be bypassed if the available capital is invested in the project studied 2. If these other opportunities carry risk similar to the project under analysis (for example, if they are other mining projects), there is no need to add incremental [ ] risk factors. On the other hand, if these opportunities are represented by the acquisition of government bonds yielding 7%, a risk premium factor must be included. The types of risks associated with a mining venture are shown in Table 1. Together they represent the "business risk" of being a mining company. The individual project risk is important to investors if it differs substantially from the business risk of the company and that of the other investments in mining ventures that we must pass up to invest in this particular project. A risk premium or discount would then be required. This can be illustrated by the following example. Consider that a company has two investment alternatives: one is the Cavender hypothetical project located in Nevada, and the other is the same project located in a civil war zone of central Europe. The second project requires a risk premium factor on the discount rate. Conversely, a project showing highly uncertain returns that are negatively correlated with the returns of the firm's other assets may decrease the company's corporate risk. Deterministic character of NPV calculation The cash flows used in NPV calculations are obtained from a set of variables. These variables are estimated using one parameter obtained from the engineering studies. Each parameter is thus a single value that is often termed a "realistic" or "best estimate" as opposed to "optimistic" or "pessimistic." However, considering that each variable can assume a range of values between the

Jan 1, 1994

By Ted C. Mathews

Pre-concentration by sorting is as old as the mining industry. Inevitably waste rock is broken with ore. Separation may take place in the stopes by selective mucking and transport, or selective loading and transport of ore and waste by 12-cu, yd. shovels and 100-ton trucks in large open pit operations. Hand sorting of 1" to 16" pieces, formerly widely practiced at underground chutes and prior to or following first stage crushing, has greatly diminished in the United States with its high per hour labor cost, although still widely practiced in other countries and, surprisingly, in some of the most modern mills where inexpensive labor is available. The economics of pre-concentration by sorting is attractive since elimination of waste at any cost less than the total milling cost will show a profit. The size at which sorting should be under- taken is determined by the size at which liberation is secured to yield a product of ore and waste, and greatly influences cost and the point in the flowsheet where sorting can be undertaken. In some instances mining costs can be reduced where an inexpensive sorting system to eliminate waste is combined with a less selective and less costly mining method. This is particularly true in deep mines where hoisting capacity determines the rate of extraction and sorting is performed underground, with waste utilized for fill. Hand sorting, while efficient, is expensive because of low sorting rates. Peale (1) quotes man hour capacity at 1/3 ton per hour on sizes 3/4" to 3" copper ore, and up to 5 tons per hour on rich iron ore sized from 4" to 16".

Jan 1, 1974

By W. W. Simpson, S. A. N. Sheya

A procedure was developed by the U.S. Bureau of Mines for recovering gold, silver, cobalt, copper, and zinc from a massive sulfide ore by a hydro-metallurgical process. The procedure consists of (1) oxygen pressure leaching; (2) iron, arsenic, and copper removal from the leach solution by precipitation; (3) selective extraction of cobalt and zinc from solution; (4) electrodeposition of the cobalt and zinc from the strip liquor; and (5) cyanidation of the pressure leach residue to recover gold and silver. The results are discussed in terms of efficiency of metal recovery and conclusions drawn as to possible benefits and broad applications of the process.

Jan 1, 1990

By R. N. Tempel, M. Lengke

Arsenic sulfide (orpiment, realgar, amorphous AsS and As2S3) oxidation rates were measured in mixed flow reactors at pH-2 and at temperatures of 25 ° to 30 °C. The measured oxidation rates for arsenic sulfide solids by dissolved oxygen are in the range of 10-11 to 10-10 mole m-2s-1 after being normalized to final surface area. The order of arsenic sulfide oxidation rates decreases at pH-2 from amorphous AsS > realgar > orpiment > amorphous As2S3. Compared to pyrite, arsenic sulfide oxidation rates are approximately 1.5 to 15 times slower than pyrite oxidation rates at pH-2.

Jan 1, 2003

By K. S. Shivakumar, R. Bhima Rao, M. R. Patil

Liberation size can be estimated using various techniques such as grain counting, frequency curves, X-ray intensity and the mesh of grind. The present investigation deals with the estimation of the liberation size of graphite flake contained in a schistose ore from Shivaganga, India. It was observed that the flakes are liberated at from 75 to 150µm. The rougher flotation concentrate obtained after grinding the run-of-mine ore for 5 min contained 54% fixed carbon. This represented the first stage of liberation. After regrinding to <90 µm in the presence of sodium silicate and guar gum, a cleaner concentrate containing 82% fixed carbon was obtained.

Jan 1, 1998

By Geoff W. Seward, Charles J. Maes

INTRODUCTION The work of the metallurgical engineer engaged in developing a leach, solvent extraction electro- winning process usually begins in the laboratory and generally proceeds via pilot plant testing before finally resulting in a commercial operation. The path leading to the development of the eventual process will involve a test program on leaching in order to derive information on the amenability of the copper-bearing material to acid dissolution, under various conditions. Commonly in North America the leach process of choice is either dump or heap leaching but it may also be in-situ leaching or the treatment of mine waste- waters and bleed streams. In the past, agitation and vat leaching have also been included in the list of options. Currently the economics of copper production dictate that only material of mineralogy and grade, capable of being leached without prior milling, will lead to a viable process. This is not necessarily true in other parts of the world where grades are higher and milling costs lower and therefore this paper con- fronts all the techniques available to the metallurgical engineer in terms of the evaluation of the solvent extraction process. Understanding the objectives of a laboratory and pilot plant evaluation of a solvent extraction route is important and these are described in some detail in this paper. Of at least equal importance is understanding limitations of such a test program and this aspect of process development is also discussed with reference to the kind of equipment and experimental practices that have been shown to produce results consistent with the objectives. WHERE TO BEGIN? At the inception of a project much information is developed on the variables of the leach process and indeed in some instances the type of leach that best suits the ore to be treated is the subject of intense investigation (Wadsworth 1983). The decision to treat the leach solution by solvent extraction focuses attention on the selection of the extractant most suited to the nature of the chemical composition of the solution. A laboratory screening program will probably need to be arranged in order to assist in narrowing the options and the following is suggested for this initial screening. Selection of a Representative Leach Solution This is preferably obtained from leaching a representative sample of the ore under conditions likely to pertain in the commercial leach circuit. As the conceptualized circuit flow sheet will undoubtedly involve the recycle of aqueous solutions via the leach, then it is desirable to know the nature of and the likely level of the impurity ions in the solution as well as the concentration of copper. The other most important parameter that is required is the free acid (Anon, Acorga Analytical Method) content of the solution. As a rule of thumb, it is always true that recovery across SX is maximized as the pH of the leach solution rises. Whilst a high pH may be desirable for SX recovery, it is not always desirable for the pH (or more correctly, the free acid) of the leach to rise too high. The working range, therefore, tends to be between 1.5-2.5 pH. At an early stage in process development it is likely that a number of ill-defined conditions will exist, making it difficult to decide on typical leach conditions in which case the screening tests should be carried out using a synthetic leach solution prepared from laboratory chemicals. In doing this the copper and free content should be carefully controlled along with the appropriate addition of major impurity ions as their sulphate salts. Preparation of Suitable Organic Solutions Manufacturers information sheets on the reagents to be tested are required at this stage The information necessary to make an informed judgment on reagent performance is: Copper Uptake Value. This will indicate the amount of copper that can be extracted by a reagent as a function of its operating concentration in a

Jan 1, 1986

By Richard Druesedow, Richard Ray

Tests at the VP-5 "C" shaft fan installation of Island Creek Coal&apos;s Virginia Pocahontas Mines were performed to determine pressure drop due to the design of the shaft collar and ductwork leading to the fan. Data from the test locations in the shaft, duct and fan indicate significant pressure losses in the system as well as significant turbulence in the flow. Data taken with a 3-D probe in the fan annulus confirmed the importance of using this type of instrument in turbulent flow applications. This data will be later supplemented with data taken after addition of turning vanes in the 45° turn adjacent to the fan inlet.

Jan 1, 1993

By K. N. (Mani) Subramanian

A 2.5 meter diameter column has been successfully used in a 100 tph flotation plant at the Harbour Lights Mine in Leonora, Western Australia, to produce a final gold bearing sulfide concentrate from the ore in a single stage. This pioneering application of column flotation cells to rougher stage was pilot tested using a 11 cm column, which demonstrated column&apos;s capability in rougher, scavenger and cleaner flotation stages. A hybrid circuit involving columns and mechanically agitated flotation cell s was adopted for the commercial installation. This paper describes pilot plant results, design of the plant flotation circuit, and early results of commercial operation.

Jan 1, 1988

By Isabel Barton

This article describes and analyzes the history of research into the origins of the uranium-vanadium(-radium) deposits of the Paradox Basin. For several decades after their discovery in 1881, research was desultory and almost entirely descriptive, fluctuating with demand for vanadium and radium. Systematic geologic descriptions, accumulated between 1910 and 1950, supported rapid progress during a government-backed surge of research in the 1950s, when uranium was the chief commodity of interest. Most of the major theoretical innovations date from this decade, including the use of concepts from solution chemistry to identify the conditions and constrain the processes of deposit formation and alteration. Research from this time was highly influential in geology at large, providing the model for sandstone-hosted uranium deposits worldwide and numerous new minerals and mineral structures. After about 1960, mine production remained high but research dwindled, and most new advances were to the details of the hydrothermal model of ore formation rather than its core concepts. By the end of the 1960s, the deposits were understood to have formed at the mixing interface of an oxidized, metal-bearing water with a reducing fluid of debated origin. Research was desultory over the ensuing decades; the source of the metals and the nature of the chemicals constituting the trap were not clearly established. They remain questions today, as do factors like the relationship between ore mineralization and the geologic history of the basin. (More in the full paper.)

Jan 1, 2025

By M. Bleeck

Calcium carbonate (CaCO3) is one of the most ubiquitous and versatile minerals found in the earth&apos;s crust. Its availability, attractive physical properties and relatively low processing cost make CaCO3 the most widely used filler material today. It is mined in three different forms - chalk, limestone and marble. Each physical form of CaCO3 has different qualities due to differences in postdepositional geology. But the chemical composition remains the same, with CaCO3 an inert component of the finished product. In the past, the paper industry largely left CaCO3 by the wayside, as it cannot withstand the acid-based papermaking process. But conversion to an alkaline system by many US mills changed this picture. Carbonate suppliers have put time and effort into research and development, demolishing barriers and creating new possibilities for what is a simple, natural product. By controlling particle size, size distribution and particle charge, the industry uses ground calcium carbonate (GCC) as a performance enhancer and as an extender for more expensive ingredients. It is estimated that the United States uses 3.6 to 4.1 Mt/a (4 to 4.5 million stpy) of CaCO3. Consolidations and mergers are taking place in the industry. Of the 12 major GCC producers in operation nine years ago, seven are left. Mineral Technology (US) is the dominant precipitated calcium carbonate (PCC) producer with more than 50 satellite plants worldwide. Other producers include Georgia Marble (French); Franklin Industries (US); OMYA (Swiss); J.M. Huber (US); ECC (English); and Filler Products (US). Global Stone PenRoc (Canada) is the only newcomer. In addition to this group, there are three small producers left in North America, each with a capacity of less than 100 kt/ a (.110,000 stpy). The trade organization operating as the Pulverized Limestone Division of the National Stone Association renamed itself the Pulverized Mineral Division, to increase its membership pool. The US paper industry is a predominant GCC con¬sumer, using approximately 800 kt/a (882,000 stpy) at an approximate cost of $130/t ($1.43/st). European paper mills pioneered alkaline papermaking. In the early 1960s, they began using GCC as filler and soon thereafter added GCC to their coating formulations. A decade later, the North American paper industry followed suit. The conversion from acid to alkaline paper production benefits the economic and performance aspects of the industry. Less pulp is needed, paper machine maintenance and effluent treatment costs are reduced, and sheet strength, opacity and brightness are increased. Perhaps most important to the reader, the sheet is desensitized to ultraviolet light, extending the paper&apos;s archival ability. CaCO3 can provide the papermaker with additional control of his paper. For example, PCC has long supplied the tobacco industry with a means to slow down the burning rate of cigarettes. Due to enhanced performance with regard to bulk and opacity, filler PCC use has risen to 1,500 kt/a (1,650 stpy) in the United States, at an approximate price of $130/t ($143/st). The majority is produced onsite at the paper mill, using "satellite plants." This concept reduces freight cost because only quicklime (CaO) is shipped to the mill, not CaCO3 slurry. The future of CaCO3 is encouraging. The amount of natural ground CaCO3 used is expected to double by the year 2005 to approximately 8 Mt (8.8 million st) worldwide. Acid papermaking practices will feel an increasing pressure to convert to an alkaline process as larger volumes of GCC containing paper enter the recycling market. CaCO3 reserves are plentiful. They will supply the ever growing demand for increasingly sophisticated paper. The plastics industry is supplied with almost 900 kt/ a (990,000 stpy) GCC at an annual growth rate of 4% to 5%. The price of a functional, inorganic filler, surface modified for the plastics industry, has an average selling price of $220/t ($243/st). GCC represents the most common filler, creating a product with higher gloss, better dielectric properties, impact resistance, weatherability and shrinkage control. CaCO3-filled plastics surround us - auto hubcaps and dashboards, shower enclosures, floor tiles, wire coatings, microwave dishes and Tupperware. The caulking and sealant industry is an enormous GCC user, with annual consumption requiring 1.13 Mt (1.2 million st) at about $44/t ($48.50/ st). Caulking and sealant may be highly filled with GCC yet undergo no adverse flow effects, with a narrow particle-size-distribution filler decreasing the binder demand. The CaCO3 industry, as well as the carpet industry, are more or less tied to the growth rates of the construction industry. It is estimated that the carpet industry uses some 680 kt/a (750,000 stpy) of GCC at about $25/t ($27.50/st). The paint industry uses approximately 300 kt/a (331,000 stpy) at a 1 % to 1.5% annual growth rate. Here, too, GCC is the dominant filler. It is used to enhance flow characteristics and color uniformity. It also extends costly titanium dioxide, creates sheen and controls roughness, hardness and tack. More CaCO3 should be used in the future as industry shifts from solvent-free or water-based formulations that can accommodate higher GCC volumes. CaCO3 is not imported or exported in any great quantity. Most areas have reserves of their own and the selling price is relatively low. Even so, quality varies from deposit to deposit. As our needs become morespecific, it remains a challenge to provide varying industries with a fitting product.

Jan 1, 1998

By Richard E. Ray

INTRODUCTION Main mine fan installations at underground coal mines must be designed to comply with the requirements outlined in 30 CFR (Code of Federal Regulations) Part 75. However, very little emphasis is typically placed on the aerodynamic optimization of shaft collar designs, ductwork configurations, and the selection of a fan isolation, or "closing door," system. Field testing has shown that poorly designed systems may produce large pressure losses between the shaft and fan inlet, as well as air turbulence problems which adversely impact fan efficiency and sometimes fan structural integrity. This paper analyzes each component of a surface exhaust fan system for an underground coal mine: • The shaft collar arrangement, • The 45 or 90-degree turning duct that isolates the fan motor from the return airstream, • Alternative ductwork arrangements for parallel fan installations, • The self-closing doors or dampers that are required for multiple fans, and • Horizontal and vertical evas6 arrangements on the discharge side of the fan. While discussion is limited to axial fan installations at shaft mines, most of the recommendations would also apply to centrifugal fan installations and to main mine fans for drift mines. FEDERAL REGULATIONS The mandatory safety standards for underground coal mines that apply to the installation of main mine fans are listed in 30 CFR Part 75.310. Stipulations which impact the design of the fan sys- tem arrangement are contained in paragraphs (a)(l), (a)(5), (a)(6),(d), and (1) of 75.31 0: • Each main mine fan shall be installed on the surface in an incombustible housing; • Each main mine fan shall be protected by weak walls or explosion doors in direct line with possible explosive forces; • Explosion doors or weak walls shall have a cross-sectional area at least equal to the area of the return air shaft or drift opening through which the pressure from an underground explosion would be relieved; • Each main mine fan shall be offset by at least 4.57 m (15 feet) from the nearest side of the mine or shaft opening; • Automatic closing doors shall be provided to prevent possible air reversals through the fans in mines ventilated by multiple main mine fans. The installation of main mine fans underground is prohibited by 30 CFR Part 75.310, (a)(l). Providing explosion doors or weak walls in line with a possible underground explosion and offsetting the fan 4.57 m (1 5 feet) from the edge of the shaft preclude the installation of the fan directly over the shaft. In addition, 30 CFR Part 75.507-1 specifies that all electrical equipment used in return air courses be permissible, which requires locating the fan motor outside of the airstream exhausted through the fan. To comply with the Part 75.310 regulations, explosion doors are typically located directly over the shaft as part of the shaft cover or &apos;bonnet." A sufficiently long section of ductwork placed between the shaft enclosure and fan, equipped with a 45 or 90- degree turn, enables the fan to be offset at least 4.57 m (1 5 feet) from the shaft and the motor to be isolated from the airstream. The need for the turn in the ductwork can be eliminated by driving the fan from the discharge side, instead of from the inlet side, and providing for a 90-degree bend in the evas6. For multiple main mine fan systems, self-closing doors can be located on the inlet or discharge side of the fan. FAN SYSTEM COMPONENTS A poorly designed exhaust fan arrangement will produce turbulent conditions at the fan inlet and significant pressure losses between the shaft and fan. Field tests have shown that pressure losses through inlet ductwork at some installations may constitute as much as 30 percent of the total mine head. Unfavorable flow conditions may cause a change in the operating point on the fan performance curve, resulting in reduced fan efficiency and pres- sure generation capability. In some instances, poor airflow distribution through the fan may create mechanical stresses that lead to fan blade or fan housing fatigue failures. The objective in designing an air duct arrangement for a fan system is to provide a low cost duct that will move the air from the shaft to the fan in a uniform air distribution and with minimum energy requirements. The optimum solution is a compromise between technical and economic factors. Aerodynamic design considerations for the shaft collar and cover, fan inlet ductwork for both single and parallel fans, self-closing doors, and evas6 are discussed in the sections below. Shaft Collar/Cover To prevent excessive turbulence and non-uniform air distribution through the fan, a smooth transition for the change in direction of airflow at the top of the shaft must be provided. However, the concrete lining of the shaft is abruptly terminated at the surface at many coal mine installations (see Figure I), creating a sharp 90-degree bend. The airflow exiting the shaft becomes heavily concentrated in the upper portion of the horizontal fan in-

Jan 1, 1997

By R. W. Smith, R. K. Mehta, X. P. Zheng, A. M. Raichur, M. Misra

Experimentation was per formed on the flotation of apatite and dolomite separately using anionic collectors in the presence of Mycobacterium phlei. Also, the flotation of mixtures of dolomite and apatite in the presence of anionic collectors and in the absence and presence of the bacterium were studied. It was found that M.phlei functions as a depressant for both minerals. However, it functions as a better depressant for dolomite than for apatite. Experimentation was extended to two different Florida dolomitic phosphate rocks. In this investigation a concentrate grade of less than 1.0% MgO was achieved with a 75% P2O5 recovery.

Jan 1, 1999

By Levent Ozdemir, Roy H. Spitzer, Don W. Deere, John B. Blyler

This paper documents the construction of the Stanley Canyon Tunnel, the geologic conditions encountered, boreability of the rock, and the impacts of the conditions encountered on the construction. The 5.05 kilometer long (3.1 miles), 3.47 meter (11.4 feet) diameter, tunnel was driven through granites with a refurbished tunnel boring machine (TBM). Difficult ground conditions encountered included shear zones, high groundwater flows, and hard, tough, and strong rock. Detailed geotechnical studies during construction revealed that microscopic changes in the rock grain boundaries due to contact metamorphism, resulted in the hard boreability. Originally scheduled for five months, the tunnel construction required nearly three years, including periods of project shutdown.

Jan 1, 1995

By Qiu Rongqing, Huang Kongxuan

This paper summarizes the results of treating arsenic­bearing refractory gold concentrates from six gold Mines of Chi­na by using bioleaching method and discusses the conditions of bioleaching processes. Results indicate that Thiobacillus fer­rooxidans T-3 Strain can intensely oxidize arsenopyrite, how­ever oxidation rate and degree differe considerably for different. During initial inoculation of 5 X 106 cell/ml. FeH contration generaly 4 -8 g/1. Sample size > 80 % - 3 20 mesh and pulp density 10 % ( w /v) and leaching time ,1 -- G clays, arsenic re­moval rate reached more than 80 % , Fe extraction ,10 - GO% desulfurization rate 30-45% after bacterial oxidation of 4-G days.

Jan 1, 1992

By Daniel P. Farjeat, Hector A. Canseco, Juan M. Schmitter

Since the early sixties, tunneling in the soft clay of Mexico City, has been done with the aid of open face shields and recently with compressed air. The use of pressurized shafts up to 1.5 kilograms per square centimeter, (147 kilo pascals) with vertical locks to launch the shields, have been used in two cases. After advancing almost 200 meters of tunnel, the same vertical locks were transformed into horizontal locks, to continue the work. Due to the high decompression times involved, it was necessary to install long locks, using the tunnel itself as part of the system. This article describes those cases.

Jan 1, 1981

By Jacques L. Mauboussin

The EUROPEAN ORGANIZATION for NUCLEAR RESEARCH (CERN) has undertaken the construction of a LARGE ELECTRON POSITRON COLLIDER (LEP), near GENEVA (Switzerland). The facility includes the main 27 km long tunnelled ring, 18 shafts, 20 experimental caverns and many auxiliary galleries. Roughly 800 000 cu.m were excavated in the molasse at 50-140 m depths, using roadheaders for caverns and auxiliary tunnels and 3 WIRTH, 4.5 m diameter, fullface TBMs for the main circular tunnel. Each TBM performs 25-30 m/day. Provisory lining behind them uses either precast concrete elements or shotcrete. Works a e being carried out by a joint-venture of European contractors led by FOUGEROLLE (France).

Jan 1, 1987

By S. R. S. Sastri, K. K. Bhattacharya, K. S. Narasimhan

Experimental results obtained on the recovery of below 0.5 x 10-3m coal fines by flotation employing geometrically similar columns of 0.056m, 0.080m, 0.100m and 0.220m diameter have been presented. The hydrodynamics of such a unit as well as its performance have been related to operating parameters to establish optimum conditions. The results have further been analysed critically in the light of the known relationship on the behaviour of similar systems to assign reasons for the critical factors established in a bid to reduce the empiricism required to be followed in the scale up of such a unit.

Jan 1, 1988