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By K. I. Cunningham, C. S. Spirakis

Two types of native sulfur deposits have been found in Lechuguilla Cave: thin late-stage crusts and massive early- formed deposits. Most, if not all, of the crusts appear to have formed by remobilization of native sulfur contained within gypsum and sulfur masses. The process is occurring today and is slowly converting the sulfur to gypsum. The absence of primary gypsum or reductants in the cave indicates that the early-formed massive sulfur deposits did not form by in-place reduction of sulfate, which is the process that formed the commercial sulfur deposits in the Delaware basin. Instead, it appears that H2S dissolved in basinal brines was oxidized by oxygen in near-surface ground waters in the cave. Fluid- inclusion data, which suggest the presence of both fresh water and saline water as the sulfur formed, support this interpretation. Although individual sulfur deposits in Lechuguilla Cave are no more than several tens of metric tons, the abundance of secondary gypsum (as much as 10 meters thick) indicates that native sulfur may once have been much more extensive.

Jan 1, 1992

By Robert C. Freas

Lime, the versatile chemical is, generally speaking, a calcined or burned form of limestone commonly known as quicklime, calcium oxide or, when water is added, calcium hydroxide or slaked lime. Almost 40 different lime products are available, a fact which has contributed to the rather loose use of the term lime as well as to confusion and misunderstanding. The term is frequently, albeit erroneously, used to denote almost any kind of calcareous material or finely ground form of limestone or dolomite. Lime is made from high calcium or high magnesium limestone, generally having a minimum of 97% combined carbonate content. Normally, high calcium lime has less than 5% MgO. When the lime is produced from a high magnesium limestone, the product is referred to as dolomitic lime. Calcination/Hydration Calcination, or the production of lime, has its origins in the earliest days of alchemy, with the general reaction class having been identified in an Arabic text printed in 1000 AD. It was much later, however, in the mid-1700s to the mid-1800s before this basic reaction became understood from scientific perspective. The pro- duction of lime has been so basic and simple that its underlying scientific principles have over the years received only intermittent investigation. Rather, much of the thought and inquiry has been directed toward the development of kilns. Thus, it has only been in the last 25 to 30 years that lime has received any concentrated scientific investigation relative to the thermodynamics and kinetics of the calcination and hydration reactions. Particular emphasis has been focused upon energy consumption and fuel efficiency. Calcination refers to a broad class of reactions, of which the limeflimestone reaction is just one, wherein a substance is heated to less than its melting point, resulting is a weight gain or weight loss. In the calcination of limestone to produce lime, the basic chemical reaction is as follows: [ ] While there is nearly universal agreement bout the equilibrium conditions related to the limestone/lime reaction above, there have been numerous and varied calcination models developed for the reaction. Recent investigations have resulted in the development of the model shown in [Fig. 1]. From this model it can be seen that calcination is a function of both temperature and CO, pressure. It does not, however, provide any indication of the rate at which the reaction takes place. Calcination is strongly time variant with different limestones. In a very broad sense this relates to the fact that the calcination reaction starts on the exterior surface of the limestone and then proceeds toward the center. As the calcination reaction takes place, the CO2 released at the interface must make its way through the lime to the exterior surface. Since calcination is limited by gas diffusion to the surface of the partially calcined limestone, the natural impurities in the stone, differences in crystallinity, grain boundary chemistry, density variations, and imperfections in the atomic lattice all play a significant role in calcination rate. Therefore, the suitability of a given limestone as a source material for lime production can be determined only after completion of adequate burn tests designed to evaluate the various limiting factors. When a coal-fired kiln system is considered, the entire process of calcination is made even more complex by the introduction of additional chemical constituents into the calcination environment. The reader is referred to the references for a more detailed description of the complete calcination reaction and the differences inherent between differing kiln systems. In the foregoing discussion it was noted that CO2 is released during the reaction. This release results in a 44% weight loss during the complete calcination of a high calcium limestone, or a 48% weight loss for a highly dolomitic limestone. The trade term for this weight percent loss is loss on ignition, or LOI. This weight loss is frequently used as a measure of the completeness of calcination. Because the calcination reaction is chemically reversible, quicklime or burned lime is frequently referred to a being highly reactive, or unstable. The more stable form of lime, hydrated lime, is commonly preferred and generally specified by the user. Hydrated lime is obtained by adding water to quicklime to produce a dry, fine powder. Quicklime's affinity for moisture is then satisfied, although it still retains a strong affinity for CO2.

Jan 1, 1994

By J. M. Hans, M. F. O’Connell, G. E. Eadie

INTRODUCTION The U.S. Environmental Protection Agency (EPA) was required, under Section 114(c) of Public Law 95-604, to provide a report to Congress identifying the location, and potential health, safety and environmental hazards of uranium mine wastes together with recommendations, if any, for a program to eliminate the hazards. The approach taken to prepare the report was to develop model active and inactive mines and locate them in a typical mining area to estimate their environmental impact. A list of uranium mines was acquired from the U.S. Department of Energy (DOE). The inactive mines were separated from the list and sorted into surface and underground categories. A literature search was conducted to obtain and consolidate available information concerning the environmental aspects of uranium mining and shortterm field surveys and studies were conducted to augment this information base. Radioactivity emission rates were measured or estimated for each mining category and were entered into computor codes to assess population exposures and subsequent health risks. The general environmental condition of inactive uranium mines was determined by walk-through surveys in several mining areas. INACTIVE SURFACE MINES We assumed that a model inactive surface mine contains a single pit with the wastes (overburden and sub-ore) stacked into a pile adjacent to the pit area. No credit for reclamation is given to the model mine. In lieu of the availability of individual mine production statistics, the model surface mine size was established from the total ore and waste production statistics for all surface mines, divided by the number of inactive surface mines. The number of inactive mines, obtained from the DOE mine listing, are summarized by type and location (Table 1). For modeling, we assumed that there are 1,250 inactive surface mines. The total or cumulative waste and ore production for inactive surface mines from 1950 to 1978 is not fully documented. Uranium mine waste and ore production statistics, on an annual basis, were available for both surface and underground producers from 1959 to 1976 (D0159-76). Annual uranium ore production for each uranium mining type are available for 1948 to 1959 (DOE79) and for combined ore production TABLE 1. Consolidated list of inactive uranium producers by State and type of mining [State Surface Underground AL 0 9 AZ 135 189 CA 13 10 CO 263 902 ID 2 4 MT 9 9 NV 9 12 NJ 0 1 NM 34 142 ND 13 0 OK 3 0 OR 2 1 SD 111 30 TX 38 0 UT 378 698 WA 13 0 WY 223 32 Total 1246 201T] for underground and surface mining from 1932 to 1942 (DO132-42). In order to estimate waste accumulated prior to 1959, the waste-to-ore ratios from the 1959 to 1976 period were plotted vs. time and line-fitted by regression analysis (Figure 1). Unfortunately, the extrapolation of the line to years prior 1959 approached zero in 1954 although surface mining began in 1950. Therefore, a waste-to-ore ratio of 8:1 was used for the period of 1950 to 1959 based on ratios estimated by Clark (C174). The waste to-ore ratios for 1976 to 1978 were estimated using the line established in Figure 1. By using waste-to-ore ratios and ore production data, the cumulative waste and ore production for both surface and underground uranium mining is estimated to 1978 (Table 2). The estimated cumulative waste from uranium surface mining for 1950 to 1978 is 1.73 x 109 MT. A crude estimate of the waste accumulated at the model inactive surface mine can be made by dividing the total waste produced to 1978 by the number of inactive mines. This, however, overestimates the waste tonnage because some of the contemporary wastes are being produced by active mines, and the waste accumulated at newer mines has increased in recent years. To adjust for this overestimate, we assumed that all mines operating in 1970 will be inactive by 1978. This eight year period is approximately one-half the lifetime of a model

Jan 1, 1981

By D. H. Carlson, J. H. Johnson, C. F. Renders

Introduction Diesel-powered vehicles are used extensively in underground mines throughout North America. The bulk of the diesel vehicles found in underground mining operations are used for loading and ore haulage, as well as for transportation of personnel and supplies. Along with the advantages of using diesels underground is the disadvantage associated with diesel-tailpipe particulate-matter emissions (DPM). The concentration of DPM in the ambient air of US underground metal mines is not now regulated by the Federal Mine Safety and Health Administration (MSHA). However, recent studies have shown DPM to be mutagenic (National Institute of Occupational Safety and Health, 1988), and the American Conference of Governmental Industrial Hygienists (ACGIH) has recommended that the exposures of per¬sonnel to DPM be limited to an 8-hr time-weighted average concentration (threshold limit value or TLV) of 0.15 mg/m3 (Anon., 1995). The authors, while making measurements in a number of US underground mines that use diesel haulage equipment, found mine air DPM concentrations ranging from 0.2 to 2.36 Mg/M3 (McCawley and Cocalis, 1986; Watts et al., 1989; Cantrell et al., 1991; Haney, 1992; US Bureau of Mines, 1992; Watts, 1992; Watts et al., 1995). If the proposed DPM TLV were to be adopted as a permissible exposure limit (PEL) for US underground mines, the proposed limit of 0.15 mg/m3 PEL would be lower than any of the concentrations measured in the earlier studies and would represent more than a 15-fold reduction from the maximum 2.36 mg/m3 concentration. A 0.15 mg/m3 PEL would also represent a 4.5-fold reduction from the average 0.68 mg/m3 measured mine ambient air DPM concentration reported in this paper. Other diesel tailpipe emissions that are now regulated underground include carbon monoxide (CO), with a PEL of 50 ppm; nitrogen dioxide (NO,), with a PEL of 5 ppm; nitric oxide (NO), with a PEL of 25 ppm; and sulfur dioxide (SO,) with a PEL of 5 ppm. Because the concentrations of these gaseous pollutants and DPM are affected by the state-of-maintenance (Waytulonis,1992), it is important that a means be developed to measure emissions from engines that are now in service to determine when maintenance is needed. The current study was the result of an inquiry by mine¬maintenance personnel who had been receiving complaints about high concentrations of diesel soot (DPM) in mine headings from load-haul-dump (LHD) vehicle operators. Mine-maintenance personnel were searching for an objective test to determine if the diesel tailpipe particulate emitted was excessive. The mine was also evaluating electronically controlled, two-cycle, naturally aspirated, direct-injection diesel engines on some of their JCI (John-Clark Inc.) load-haul-dump (LHD) vehicles. These LHD vehicles were used to haul freshly blasted ore from mine headings to a feeder breaker. The feeder breaker breaks down the larger chunks and feeds the broken ore onto a conveyor. Michigan Technological University, in past studies, developed an emissions-measurement apparatus (EMA) ca¬pable of measuring diesel vehicle tailpipe pollutant concentrations (Chan et al., 1992; Chan et al., 1993; Carlson et al., 1994). At the time of the study reported here, most of the mine's LHD vehicles used a 12-cylinder, four-cycle, naturally aspirated prechamber diesel engine. The study was undertaken in cooperation with mine maintenance supervisors from late 1992 through July 1993. The objectives were to compare diesel exhaust emissions between the 6-cylinder, two-cycle, electronically controlled, direct-injected diesel engine and the 12-cylinder, four-cycle, prechamber diesel engine and to, then, use the data collected, in conjunction with mine ambient air measurements, to demonstrate the application of the "deterioration factor" (Chan et al., 1992), which is a measure of the state-of-maintenance of mine-vehicle engines that are now in service. The information would be used to identify vehicles that need maintenance to reduce emissions. The data reported here are unique in the sense that they combine underground diesel vehicle ambient and tailpipe

Jan 1, 1999

By N. D. Jagger, I. M. Arbuthnot

Introduction Over the last five years, a large number of small- to medium-sized carbon-in-pulp treatment plants have been built in Australia, most designed to treat between 250,000 t/a and 1.5 Mt/a of ore. Because of the limited capital resources and tight cash-flow positions of these relatively small mining companies, the primary requirement was often to get a plant built and operating in a short period of time and at minimal capital cost. Therefore, since the inclusion of both pre-leach and tailings thickeners represents an obvious and significant capital cost, most of these plants were built without thickeners or even detailed, cost-benefit analyses on their inclusion. In some cases, the increasing use of High Rate Thickeners (HRTs) in the mineral processing industries has, however, resulted in a reassessment, because of their considerably lower cost. This reassessment was triggered primarily by the need to conserve water in arid mining areas where borefields are costly to install and water is limited. With the startup and operation of these installations, the resulting significant savings in cyanide consumption has been recognized, in many situations, as a primary justification for the installation of HRTs. Solution balances Degradation of cyanide occurs in the tailings water discharge to slime dams. The degree of degradation (cyanide loss) in the water recovered depends on a number of factors, but it is usually assumed to be about 90%. The most important mechanisms of CN loss are through HCN losses and oxidation by oxygen in the air, which also assists in the hydrolysis of CN. These mechanisms are supported by the large dam surface area and the long retention time of the tailings water in the dam. By thickening the CIP tailings at the plant and recovering as much tailings water as immediately possible, these losses are avoided. The retention time in an HRT is less than three hours, and the surface area is relatively small. Therefore, CN losses are negligible, which is not necessarily the case in conventionally-sized thickeners. Fig. I shows a block diagram of a 100-t/h gold plant without a thickener. In this example, 50% of the tailings water pumped to the tailings dam is recovered, and the CN concentration of the returned water is 10% of the tailings CN concentration of 150 ppm. Fig. 2 represents a plant with an HRT on tailings, thickening to 55% w/w solids. In this case, the thickened tailings are pumped to the dam, and 25% of the contained water is recovered. The recovered water and the mill make-up water are not sent directly to the mill; instead, they are added to the thickener feed and mixed with it prior to thickening. By doing this, the tailings are effectively washed, and the additional cyanide is recovered. Solution balances over these two circuits show cyanide recoveries of 5% and 65%, respectively. Thus, the thickener use increases cyanide recovery by 60%. Fig. 3 shows a two-stage HRT circuit in a countercurrent decantation (CCD) configuration. In this example, an additional 13% of cyanide is recovered through the use of the second-stage unit. This configuration can be justified when residual cyanide levels are high. Capital and operating costs - Case study 1 Illustrative cost figures are based on a CIP tailings thickener installed, in early 1988, as part of Dominion Mining's treatment plant at Paddy's Flat near Meekatharra. (All costs within this paper, unless stated otherwise, are in Australian dollars.) [Assumptions: Feed rate 150 t/h Ore moisture content 5% Leach density 40% solids Underflow density 55% solids Residual cyanide in tailings 150 g/m3 Flocculant dosage 15 g/t Consumable costs: Water $ 0.40/m3 Cyanide$ 2.00/kg Flocculant$ 4.50/kg Power$ 0.12/kWh Length of tailings pipeline2,500 m Capital costs of the major items involved in the thickener installation are given below: Thickener unit, 15-m diam $ 230,000 Water return pumps 12,000 Water recirculation pumps 28,000 Feed box 5,000 Flocculant make-up system 18,000 Flocculant storage tank and dosing pumps 11,000 Piping and valves 85,000 Electrics and instruments 47,000 Civil works 29,000 Installation 69,000 Total cost$ 534,000] Savings in capital costs that can be attributed to the

Jan 1, 1993

By R. K. Mehta, C. W. Schultz, J. B. Bates

Introduction The Mineral Resources Institute, The University of Alabama, has for the past three years been engaged in a program to develop a beneficiation system for eastern (Devonian) oil shales. One objective of that program was to evaluate advanced technologies for effecting a kerogen-mineral matter separation. Column flotation was among the advanced technologies selected for evaluation. Early in the program it was shown that column flotation was superior to conventional (mechanical) flotation and to the other advanced technologies being evaluated. The investigation then proceeded toward the further objective of defining the optimum operating conditions for column flotation. One observation made in the course of optimization testing was that introducing the feed into the froth (above the pulp-froth interface) resulted in an improved combination of concentrate grade and kerogen recovery. This observation was reported in a previous paper (Schultz and Bates, 1989). Because the practice of maintaining the pulp froth interface below the feed point is contrary to "conventional" practice, it was decided to subject the observation to a systematic series of tests. This paper describes a recent series of tests and the results that were obtained. Experimental equipment and procedure The arrangement of the column cell and auxiliary equipment for continuous flow testing is shown schematically in Fig. 1. The feed sump [O] is filled with a sufficient volume of prepared sample to permit a large number of tests to be performed (typically 12). Past experience has shown this is necessary to control sample variability and variability in the size distribution resulting from ultra fine grinding. The feed slurry is maintained at about 20% solids and is constantly recirculated and stirred. The sample is metered from the circulating pipe by a peristaltic pump [O]. The feed slurry is diluted with reagentized water [O] by a second peristaltic pump [O]. Wash water [O], also reagentized, is supplied through a third peristaltic pump [O]. While this feed system may seem unduly complex, it does permit users to independently vary either the wash water rate or the net solids content of the cell. In the tests reported here, the feed rate and net percent solids were constant at 12.5 gms/min. and 3.3%, respectively. Diluted feed enters the column through 6.35 mm-diam (0.25 in.-diam) copper tubing and is discharged upwardly at the center of the column. Tailings are discharged through flexible tubing that can be adjusted so as to control the position of the pulp-froth interface. The column is 76.2 mm-internal-diam (3 in.-internal-diam) and 1090 mm (43 in.) high. It is made from lucite tubing and is fitted with a 51-mm-diam (2-in.-diam) fritted glass air sparger having an average pore diameter of 50 µm. In performing a series of tests, the concentrate and tailing are allowed to discharge continuously. The system is allowed to equilibrate for 30 minutes after the pulp and froth reach operating levels. Concentrate and tailing samples are taken simultaneously for timed intervals (five to 15 minutes, depending on the volume of sample desired). After sampling, a change in operating conditions is made and the system is again allowed to equilibrate. The tests to determine the effect of the pulp-froth interface level were part of a larger series of tests in which the objective was to optimize the conditions for a rougher flotation stage in a two stage circuit. The sample used in this series of tests was an Alabama shale ground to d90 = 23.1 µm and d50 = 7.9 µm. The operating conditions remaining constant in this series of tests were as follows: Column height - 1600 mm (63 in.) Air sparser - 50 µm (average pore diameter) Spray water - 130 cc/min. Feed rate - 12.5 gm/min (0.4 oz per min) (dry solids) Percent solids - 3.3% Frother (Dowfroth 250) - 45 ppm The variable test conditions are tabulated in Table 1. Positions of the pulp level (pulp froth interface) and feed entry are presented as a percentage of column height (as measured from the face of the air sparser). These test conditions are presented Fig. 2. At each of these test conditions, individual tests were performed at varying air

Jan 1, 1992

By Jozef S. Stachulak

Inco Limited operates 11 underground mines in the Sudbury District. The mines are located on the rim of the Sudbury Basin, an oval with the axis in the range of 27 and 60 km. The ore dips to at least 3000 m below surface. The ores are mined primarily for nickel and copper. Total ore production from underground is in excess of 55,000 tons per day. Over 40 surface fans have been installed since the late 1960's. All of the fans are adjustable pitch, axial flow units. A major factor influencing ventilation design in the last 30 years has been the introduction of diesel equipment underground. Volumes per fan have ranged from 60 to 330 (cubic metres per second), with motors from 100 to 2500 hp. Fans of the axial flow type have been in common use for main fan installations at Canadian mines for many years. The standard arrangement has been to mount these fans horizontally, i.e. with the fan shaft and the long axis of the housing horizontal. This is a natural arrangement for an underground fan, but for a surface installation, a vertically mounted fan has definite advantages. The surface area taken up by a typical vertical fan installation is generally about one quarter of that with a horizontal fan of the same capacity. (1) This is not a problem with isolated fans and flat surface outcrop sites, but where the installation is to be near existing buildings, or where there are poor surface soil conditions, space and cost considerations greatly favour vertical fans. MAIN FANS INSTALLED There are 43 main surface fans in operation at 11 mines. Twenty-seven of these fans are supply units, and 75% of them are vertical installation. The remaining 16 units are main surface exhaust fans, with predominantly horizontal installation. Within the last five years, some 20 main booster fans have been installed underground at several mines. Axial flow fans, with adjustable pitch blades, are used for both surface and underground installations. Exhaust fans are equipped with stainless steel or cast aluminum blades. Main underground fans are arranged horizontally and the majority of them have a floating shaft between the fan shaft and the motor. (2) The size of the fans in service varies from 1.8 m to 5 m in diameter, with the majority ranging from 1.8 to 2.5 m. The pressure produced by these fans varies from 0.25 kPa to 2.0 kPa. At Inco Limited, two main fans in parallel are preferred, rather than a single fan, so that if one fan fails, the remaining fan can still supply up to 70% of normal air quantities, while the damaged fan is repaired. This requires closure doors on each of the fans so the fan can be isolated in case of failure. It is more expensive than a single fan, but results in less production interruptions. The fan installations are well away from sharp inlet and outlet bends. FAN DESIGN INTEGRITY Both the mine operator and the fan manufacturer must understand that the main fan is critical to the mine operation, and that everything technically possible must be done in design and manufacture to ensure the highest degree of reliability. Some of the design parameters and criteria, based on Inco experience, are outlined and discussed below. RESONANT FREQUENCIES AND HOUSING MODEL ANALYSIS Any fan assembly will have many different resonant frequencies. It is a challenge to the designer to arrive at a design in which forcing frequencies do not coincide with any of these resonant frequencies to produce unacceptable vibration levels in operation. Finite element analysis is a useful tool that can be used to identify the most critical of these, so that housing and blade stiffness can be adjusted to change any resonances that might be close to forcing frequencies. Shaft critical speeds should be at least 25% above the fan operating speed, and there should be sufficient separation between other resonant and forcing frequencies to avoid excitation that might result in high vibration levels. QUALITY ASSURANCE Radiographic Blade Examination Mine fan blades are normally cast in small lots. To ensure that the castings are sound, a full radiographic examination is recommended in the highly stressed lower 1/3 of the blade.

Jan 1, 1995

By K. Mulukhov

Introduction The depth of open-pit mines continues to increase. The typical designed depth for large quarries is now in the 300 to 600 m (1,000 to 2,000 ft) range, with some quarries even reaching depths of 700 to 800 m (2,300 to 2,600 ft). The cost of handling material from such deep mines can be 60% to 70% of the total cost. The capacity of using a cyclic means of transport, such as trucks or skip hoists, continues to drop as the depth of the mine increases. In addition to the problem of maintaining relatively low slope angles for haul roads, increases in environment pollution and increased fuel consumption make the use of trucks ineffective below 100 to 150 m (300 to 500 ft). In handling large quantities of bulky materials, i.e., more than 15 to 20 Mt/a (16 to 22 million stpy), under fixed terminal points, belt conveyors are preferable in almost in all cases. An exception is when the lump size of the load is too high for the conveyor to handle. Belt conveyors in mines can handle loads having a maximum lump size of no more than 250 to 300 mm (10 to 12 in.). The use of a movable crusher prior to the belt conveyor in an open-pit mine would make the process very expensive. Such an operation could not compete with truck haulage on a cost basis. Belt-car conveyor The main factor that restricts the lump size for conventional belt conveyors is collision impacts between the load-carrying belt and idlers. These impacts are completely eliminated in the belt-car conveyor, which was previously described by Mulukhov (1977). The unique feature of the conveyor is that the load-carrying belt is supported by moving cars. Each car consists of a troughed cross strap that supports the belt with the load and two rollers, on which the cars move along upper and lower rails. Cars are connected by two endless chains revolving over sprockets. The system uses conventional drive pulleys, tail pulleys and return idlers. There are no fasteners between the carrying belt and the car cross straps. The conveyor belt simply drives the cars along by the force of friction between the belt and the straps. A belt-car conveyor was first installed in 1970 at the Karatau mining operation in Kazakhstan. The system was an elevating conveyor with an inclination of 20° and was located on the final part of the truck haul road. The conveyor transferred blasting phosphate ore from the quarry to the crushing plant. A bunker was mounted between the trucks and the conveyor. The maximum lump size was typically 1,200 mm (48 in.). However, there were some lumps as large as 1,500mm (60 in.), which was greater than even the belt width (Fig. 1). Because the prototype belt-car conveyor was intended only for industrial trails, it had a relatively short length of about 50 m (160 ft). Nevertheless, the conveyor yielded a substantial profit over the cost of truck haulage. Naturally, the first question one might ask is why such a system is not now widely used. The reason was the further development of a huge transporting complex for the whole quarry at the base of the belt-car conveyor. After successful industrial trials from 1971 through 1973, which demonstrated the systems engineering feasibility, the State Planning Committee of the USSR included the development of the belt-car conveyor in its plan for the development of the national economy. However, mistakes were made. First, a single organization was not in charge of the development. Second, a plant that was not specialized in belt conveyors was put in charge of the design and manufacturing of the conveyor system. In addition, the crusher was installed before the stacker, which negated the major advantages of the belt-car conveyor. For loading the conveyor, a special blade feeder with pendulum suspension was developed, but it was found unreliable. The main conclusion was that, in-

Jan 1, 2003

By Don Rose

I. INTRODUCTION The Atlanta Research Chamber is an UMTA-sponsored applied research facility studying tunnel design, excavation and support systems. It is located directly below Peachtree Street in downtown Atlanta, Georgia. It is included in the CN-120 contract to build the Peachtree Center subway station, a part of the Metropolitan Atlanta Rapid Transit Authority (MARTA) multi-billion dollar "Phase A" 53-mile transit project, which is 80 percent funded by UMTA with 20 percent provided by local authorities in Atlanta and Fulton and DeKalb counties (Figure 1). The Research Chamber is located immediately south of the main Peachtree Center Station cavern, and is in part an enlargement of the pre-existing Pilot Tunnel excavated to provide bidders on the CN-120 contract a view of the rock conditions (Figure 2). The Research Chamber lies some 3 meters above, and is approximately parallel to, the twin Running Tunnels. The Research Chamber is 18 meters long and has the same 5.5 meter diameter horseshoe shape design dimensions as the twin Running Tunnels.

Jan 1, 1979

By Herbert P. Haschke

I. Changes made by the Economic Recovery Tax Act of 1981 (ERTA) which have an effect on investing in tax shelter investments A. Maximum ordinary income tax rate on individuals is now 50%. B. Maximum tax rate on long-term capital gains for individuals is now 20%. C. Maximum tax rate on alternative minimum taxable income for individuals is now 20%. D. Although not changed, the "add-on" 15% minimum tax on tax preference items and the possibility of paying the alternative minimum tax cause individual U.S. income tax planning to continue to be very complex. II. Real Estate A. The new Accelerated Cost Recovery System of depreciation (ACRS). 1. Mandatory for assets placed in service after 1980. ACRS will eliminate most if not all controversies with the IRS over depreciation. 2. Real property now depreciated over 15 years. Component depreciation is not available. 3. Accelerated depreciation is available - a 175% declining balance method with an automatic shift to straight line method at the optimum point in the asset's life. Alternatively, a taxpayer can elect to use the straight line method.

Jan 1, 1982

By Paul D. Tavares, John Forbes, Eric R. Drooff

As the existing major U.S. - Canada rail tunnel was too small to accommodate double-stack container cars, Canadian National Railways is constructing a new parallel tunnel with a 50% greater diameter. Starting in Sarnia, Ontario, the new shallow-cover bored tunnel runs under a large research building and several other structures in an oil refinery before travelling under the St. Clair River. Maximum predicted settlements to be caused by the 9.2 meter diameter earth pressure balance TBM in the soft St. Clair till were in excess of 100 millimeters. As such, these settlements had the potential to cause severe damage and disruption to the existing refinery. In order to reinforce the soft clays below the refinery research building and other structures, and to compensate for settlement caused by the tunnelling, the ground improvement method of Soil Fracture grouting was successfully applied for the first time in North America. This paper will describe the details of both the soil fracture grouting application and the elaborate electrolevel remote monitoring system used.

Jan 1, 1995

By Neal Rackstraw, Mamo Meaza

Washington Metropolitan Transit Authority Section E8a consists principally of twin tunnels 1,713 meters (5,620 ft) long, excavated to a diameter of 6.35 meters (20.83 ft) using an open face shield, with a primary lining of four unbolted precast concrete segments expanded at the crown, and a full circle unreinforced concrete final lining with a finished tunnel diameter of 5.08 meters (16.67 ft). The tunnels, constructed in Cretaceous age sediments, are mostly in stiff to hard clay except for a few areas along the alignment that contain dense to very dense fine sand in the upper portion of the tunnel that created a mixed face condition. At two locations unexpected heavy water inflows produced very serious problems in both tunnels. Water flows in the interface area were generally more than 378.5 1pm (100 gpm) and reached as high as 2,649.5 1pm (700 gpm) in the mixed face ground. The nature of the tunnel inflows, the consequences, and the corrective measures adopted will be the subject of this paper.

Jan 1, 1991

By Leonhard Schmid

INTRODUCTION In the region of Zurich several of the Swiss Interstate Highways meet and have to be linked to form a continuous network. The subalpine glacial topography with hills, ridges and lakes requires the construction of many tunnels (Figure 1). The Milchbuck Tunnel forms part of the inner arterial highway system of the city of Zurich. It passes underneath the ridge separating the Limmat and Glatt valleys. In its final configuration the tunnel will consist of two parallel three-lane tunnels (Figure 2). For the time being only one tunnel is constructed. This tunnel will be operated with two outbound (climbing) lanes and a single inbound (descending) lane (Figure 3). The owner, the Canton of Zurich, awarded the contract for the construction to a joint venture with the following three contractors: Locher & Cie AG (leader), Prader AG, A. Brunner's Erben, all located in Zurich. The actual freezing was performed by subcontractor Philipp Holzmann AG, Mannheim, West Germany. Electrowatt Engineering Services Ltd, Zurich, was design engineer and performed the construction supervision.

Jan 1, 1981

By Wiston Rocher

The Minera Ojos Del Salado (MINOSAL) concentrator is located 20 km (12 miles) southeast of Copiapo, Chile. It processes copper ore provided by its Santos and Resguardo mines at a capacity of 4.2 kt/d (4620 stpd). In 1982, Phelps Dodge acquired the majority interest in MINOSAL to explore the property's potential. At that time, the concentrator plant was a small, antiquated installation with a capacity of 674 t/d (740 stpd). At the time of the pur¬chase, the majority of the ore was produced from the Santos and Resguardo mines, but custom ores of copper, silver and gold were also treated. The acquisition occurred at a crucial time in the copper business, since the price had fallen to nearly 50 cents/ lb. In 1985, Phelps Dodge purchased the outstanding interest in MINOSAL and acquired 100% of the property. Today, after numerous improvements and two expansions, the concentrator is a combination of old and new equipment. The regrind mills, for example, have been in service for about 50 years, while the ceramic concentrate filter represents the latest technology.

Jan 1, 1993

By Claus Becker

At the RETC 1979 in Atlanta/Georgia it was reported about the beginnings of the Hydroshield system and the first site experiences. Since that time over 10 miles of tunnels in various diameters have been built with that system, and the discussion of the two most interesting projects will be the main topics of this paper. AURELIA TUNNEL, ROME In Rome, the railway system is going to be completed and modernized. Therefore, it is necessary to tunnel through the Roman hills. Part of the new railway system is the Aurelia Tunnel, located some 100 feet below the famous Via Aurelia Antigua, close to the Vatican. The double track railway tunnel has an internal diameter of 9.44 m and a length of 2.410 m. It runs through completely differing soil conditions. (Fig. 1) The first 1000 t are situated in watertight clay, the rear part in water bearing loamy, clayey sands with interlayers of loamy clays. The figure shows the soft change-over from clay to sand. The groundwater head is between [ ]

Jan 1, 1983

By W. Stark

Many developing countries with little or no previous environmental regulation are now promulgating air quality regulations which require facilities to demonstrate compliance with both process parameter based emission limits and ambient air quality standards. For many companies with operations in developing countries, an air quality compliance program at the facility level is lacking and these companies have very little background information to work with in developing a program for assessing and maintaining compliance. This paper will present a methodology for development of a comprehensive air quality compliance program for operations located in developing countries with new environmental regulations. The experiences of the authors in developing an air quality compliance program for a mining operation located in South America will be used as a basis. A program for meeting new air quality requirements can be developed in seven steps. The seven steps are: (1) regulatory review and project scope; (2) design and implementation of a baseline air quality monitoring program; (3) site visit, (4) development of a complete air emission inventory; (5) dispersion modeling; (6) compliance analysis and compliance plan; and (7) project completion.

Jan 1, 1998

By Silvano Perla

Peabody Western Coal Co. operates two mines in northern Arizona. The Black Mesa and Kayenta coal mines are unique for several reasons. They both have only two customers: Southern California Edison’s Mohave generating station near Laughlin, NV and the Navajo generating station near Page, AZ. Both mines are close together. Between the two, they ship nearly 11 Mt/a (12 million stpy) of coal. Material is shipped to the power stations in two ways. This first is a 440- km-long (273-mile-long) underground slurry pipeline, the longest in the world. The second is an 11 -km-long (17-mile- long) long conveyor to four storage silos. Product is then shipped by train from the storage silos to the plant sites. Another factor makes these mines unusual. They are both situated in the middle of the 68,795km2 (17 million- acre) Navajo Reservation, home to the Navajo Nation and Hopi Tribe members. Nearly 90% of the Black Mesa and Kayentamines’ 850 employees are tribe members. Peabody Western Coal is the United States’ largest private employer of Native Americans.

Jan 1, 1995

By Ivano Andreis, Giovanni Valent

The Yindaruqin Irrigation Project is designed to irrigate 57,000 ha of land in the central part of Gansu Province of People's Republic of China. Water for the Project will be diverted through an 87 Km long feeder canal which will convey 36 cubic meters per second from an intake on the Datong River. (Fig. 1) The feeder canal consists of approximately 10 Km of open channels, 75 Km of tunnels and 2 Km of siphons and aqueducts. This is the biggest self-flowing irrigation water control project in the Northwest of China and it is among the most interesting projects in China for the geological and environmental conditions. Nearly half of the investment in the Project is a World Bank Loan and contractors were chosen by means of international bidding. The Italian Company, CMC Ravenna, succeeded in answering the bidding for the construction of tunnels 30A and 38, perhaps the most difficult in the Project, submitting a TBM alternative to the conforming drill and blast contract recommendation.

Jan 1, 1993

By Leonard W. Postregna

Prior to January, 1986, the city of Austin, Texas was on notice by the Federal Government's Environmental Protection Agency and the Texas Water Commission, that they were illegally polluting the Colorado River. The wastewater facilities presently in use in South Austin were inadequate to handle the vastly increasing population. The city of Austin, Texas hired Parsons Brinckerhoff Quade and Douglas, Inc. and Dannenbaum Engineering Corporation to design and supervise the construction of an update to their existing system to alleviate this problem. A major criterion handed down to the Designer and Contractor was that the system had to be completed within a two year time frame. The biggest link to the entire design was a tunnel project consisting of 43,000 linear feet (13,106 m) of tunnel, eight feet (2.44 m) in diameter with connecting structures tying into existing and future facilities. The project was awarded to S. A. Healy Construction Company of McCook, Illinois with a low bid of $24,755,000.00. This article summarizes the techniques used in the construction of this tunnel project.

Jan 1, 1989

By R. H. Sailors

During the last 15 years, the North American Mining Industry has experienced increasing usage of cast high chromium white iron grinding media in wet grinding applications. First introduced into the cement industry (klinker grinding) where wear rate reductions of four or five times have been realized compared to steel balls, this same grinding media serve the iron, copper, and phosphate markets by resisting not only impacts and abrasion, but also corrosion which occurs in wet grinding. In these markets, single mills containing 250,000 kg of grinding balls are not un-common and hence the economic potential with respect to lower wear rates can be significant. Add to reduced wear rate the increased production efficiency that is possible and it is readily understood why much of the mining industry uses cast high chromium grinding media. Metallurgically, the alloys used consist of a martensitic matrix containing hard M7C3 eutectic carbides. Carbon (2.0 to 3.5%) and chromium (12 to 277.) are characteristically used to arrive at the desired properties.

Jan 1, 1989