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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 Randy Vranes

Sophisticated information technology, quality management systems and a commitment to the environment has allowed Solutia’s Soda Springs phosphate mine and elemental phosphorus plant to reduce operating costs, improve product quality and contribute to the sustainable development of Idaho’s industrial minerals. Soda Springs background The Soda Springs phosphate mine and plant is a joint venture between Monsanto and its operating partner, Solutia Inc. The operation is based in southeastern Idaho in the Western Phosphate Field. The plant is located about 275 km (170 miles) northeast of Salt Lake City, UT and about 4 km (2.5 miles) north of the town of Soda Springs.

Jan 1, 1999

By John Kinneen, Sunghoon Choi, Arne Fareth, Edward S. Barboe, Kyle R. Ott

A new water treatment plant is to be constructed for a portion of the water supply system of New York City. The project includes construction of new water conveyance tunnels crossing perpendicular and sub-parallel to the existing pressurized water tunnel at two locations. A study was conducted to evaluate the impact of new tunnel construction on stability of the existing tunnel and hydrojacking potential in the surrounding rock mass. This paper introduces a practical approach to evaluate tunnel interaction, expressed in terms of rock-liner interaction, stress-induced stability and hydrojacking failures. A simple method to evaluate hydraulic and mechanical properties of concrete liners and jointed rock mass is also presented.

Jan 1, 2007

By Louis W. Cope

Mongolia is known in western countries as the location of rich dinosaur fossil fields. However, the country is even richer in mineral resources. The Ministry of Industry and Trade reports 6,000 mineral occurrences of 80 metals and minerals. Mongolia has already attracted much foreign investment in the mining sector. Mining is the country’s largest industry. It is responsible for 17 percent of the gross domestic product, 65 percent of industrial output and 57 percent of export earnings.

Jan 1, 2006

By Craig Bournes, David Chapman, Michael DiPonio

Historically, boulders are a frequent source of problems in soft ground tunneling. During tunnel construction, the need to manually break and remove boulders as obstructions causes delays to the project. Repairs to a tunnel boring machine (TBM) can also be a source of delays. Managing these problems is difficult since normal soil investigation techniques do not accurately predict the presence or frequency of boulders. This has lead to considerable number of claims for extra costs and delays during the construction of soft ground tunneling projects. These issues are exacerbated in pressurized face tunneling systems where there is limited access to the TBM cutterhead for obstruction removal and/or cutterhead maintenance. The project team that built the Big Walnut Augmentation/Rickenbacker Interceptor (BWARI) Project for the City of Columbus successfully managed the design and construction of a 4,267 mm (14 ft) diameter soft ground tunnel with high boulder concentrations in a complex geologic setting. The management of the boulder issue tookplace during several phases of the project.

Jan 1, 2007

By James J. Gusek

The Atlantic City iron mine is located in Fremont County in central Wyoming, near the historic gold mining camps of Atlantic City and South Pass City. The mine lies on the southeast flank of the Wind River Mountains at an elevation of about 2,470 m (8,100 ft) above sea level (Fig. 1). An 81-h M2 (200-acre) tailings impoundment at the Atlantic City taconite operation has been a source of blowing dust. The site qualified for reclamation under Wyoming&apos;s Abandoned Mine Land (AML) reclamation program. The ultimate reclamation construction incorporated commercially available organic amendments and fertilizers into a 300-mm- (12-in.-) thick cap of a sterile clay loam cover material. The water level was lowered in an adjacent flooded mine pit that was feeding uncontrolled seeps into the tailings basin. A wide tailings pond spillway was built that allows flood control while minimizing seasonal water-level fluctuations in the pond. Rock berms and rock snow fences were built for wind-erosion control and soil-moisture preservation. Future work includes tree planting in the protective wind/ snow shadows of rock berms and rock snow fences and plugging a decant pipe. US Steel opened the Atlantic City iron mine in August 1962. It closed in 1983. Taconite iron ore was mined by surface methods. The ore was magnetically separated from the waste. The magnetic concentrate was mixed with bentonite, formed into pellets and roasted to produce a blast furnace feed stock for US Steel&apos;s blast furnaces in Geneva, UT. The pellets were hauled by rail to the Geneva Iron Works. The Atlantic City Mine produced three types of wastes. These included waste rock overburden from mining, a coarse-grained (sand to small-gravel sized), nonmagnetic reject rock termed "cobbs" and fine-grained tailings ranging from 600 to 75 µm (28 to <200 mesh). The waste rock has been used as an aggregate/rip rap source by a local contractor and the Wyoming Department of Transportation (WYDOT). WYDOT also

Jan 1, 1998

By M. Lonoff

G. Campbell The paper by M. Lonoff looks at the importance of byproduct prices on copper production. The paper develops several interesting points on this topic, but there are some points in the theoretical discussion that could use further development. A couple of these theoretical topics will be considered here using the paper&apos;s framework of assumptions without concern about how realistic the assumptions are for the copper industry. Lonoff&apos;s theoretical model is based on a static, homogenous world where all copper/byproduct deposits are identical, and these deposits are the main source of these metals. Due to the importance of the byproducts to the production decision, the term coproducts will be substituted for the term byproducts to reflect more accurately this condition. The conclusion is drawn that a rise in coproduct prices necessitates a decrease in copper prices because copper production is increased as a consequence of increased coproduct production. The observation that real world behavior does not always follow this pattern is dismissed as "speculation." However, it can be easily shown that within this simple framework market forces might create the observed behavior. The key question to be answered is what causes coproduct prices to rise. Three possible cases will be used as illustrations. The first case is the one implicity assumed by Lonoff. Here, the demand for coproducts increases at all prices (an outward parallel shift of the demand curve) with the demand for copper remaining constant. As reported in the paper, the results are higher coproduct prices, an increase in coproduct production, and lower copper prices due to the resulting increase in copper production. The second case is when the supply of coproducts are constrained-leading to higher market prices. Copper production, as required by the simple model, is also reduced which Reply by M. Lonoff Campbell&apos;s comments give me an opportunity to clarify a few points in "Copper and Its Byproducts." That paper examined the affect of byproducts on the copper market and considered the relative riskiness of multi-metal deposits from an investment standpoint. That brief note made no attempt to be exhaustive. Nevertheless, Campbell&apos;s comments miss the point of the paper and necessitate a response. My "theoretical model" was static and I freely conceded after introducing the assumption of identical deposits that "none of the assumptions hold." These simplifying assumptions allowed me to bound the possible effects of higher byproduct prices on the copper price by assuming identical deposits and computing average price effects. I took great pains to subsequently explain that since deposits are not identical the identity of the marginal deposit might change due to byproduct price changes. In this case, average price affects will overstate the affect of byproducts on copper prices. According to Campbell, I dismiss as speculation any "real world behavior that does not always follow this pattern" (of metals prices moving in opposite directions). In the byproducts paper I noted that gold, copper, and silver price movements often parallel each other on an intra-day or day-to-day basis. Such movements are largely due to market anticipation of or reaction to changes in interest rates that affect carrying costs of metals stocks. Other movements may be due to anticipated changes in available supply or physical requirements. I loosely grouped these effects as speculative. To distinguish this statement from the main thrust of the paper, I contined "In the long run if gold and silver prices remain high the equilibrium copper price will be lower." The paper set out long run equilibrium conditions rather than attempting to explain price movements on a daily basis or on a year-to-year basis.

Jan 11, 1983

By J. MCGUIRE, E. J. Haas, B. Demich

Near misses can enhance awareness of the potential causes of injury and prompt safety management initiatives. Although most companies require near-miss reporting, the value of these reports and how they influence subsequent actions or controls to reduce on-the-job risks are unclear. Researchers at the U.S. National Institute for Occupational Safety and Health (NIOSH) conducted a case study with an aggregates company in which near-miss reports were analyzed at each of their locations over an entire quarter during the summer of 2018. Within that quarter, workers recorded 249 near misses. The analysis provides insight into ways that risk-communication and management programs can be improved to reengage workers and their situational awareness on the job.

By Debasis Deb

The first version of the longwall strata control and maintenance system (LSCMS) has been successfully developed by the Strata Control Re¬search Team at the University of Alabama. The system includes a software package designed to monitor shield leg pressure, interpret the data and predict the forthcoming roof loads and the existence of geological anomalies. It can also detect maintenance problems of the face equipment. The LSCMS was installed underground and connected to the surface terminals at Drummond&apos;s Shoal Creek Mine in Alabama. The system has proven to be a useful and practical system for improving safety and efficiency in coal-mining operation. The Shoal Creek Mine is a new, large coal mine operated by Drummond with two longwall panels. The Blue Creek seam is the primary source of coal production. It is located at a depth of 275 to 440 m (900 to 1,450 ft) below the surface with an approximate 2% southwest dip. Typical strata conditions are represented by two stratigraphic columns (Figs. la and b). A fireclay/dark shale parting (called Middleman) 0.2- 0.7-m- (1- to 2.5-ft-) thick exists between the Mary Lee and Blue Creek seams. Mining conditions vary from area to area. This mainly depends on the Blue Creek seam thickness that ranges from 0.9 to 2.9 m (3 to 9.5 ft). The thickness of the Mary Lee and the parting is relatively consistent with a to¬tal of 0.9 m (3 ft). They are mined when the Blue Creek seam becomes thinner. In most of the panels, the immediate roof is composed of sandy shale 1.5-

Jan 1, 1997

By C. Laughton

The Deep Underground Science and Engineering Laboratory (DUSEL) is to be constructed at the defunct Homestake Mine, in South Dakota. As a cornerstone of DUSEL?s research program, it is intended to conduct experiments into the fundamental nature of neutrino and proton particles. The experimental apparatus required to con-duct these experiments will be housed in large rock cavern(s) sited at depths of over a kilometer. As conceived, these cavern(s) are planned to have mined volumes of up to half a million cubic meters, and spans of approximately 50 meters. The paper will describe key end-user requirements, and outline the in situ rock conditions and mine site constraints that are likely to play determinant roles in identifying candidate sites and developing cost-effective design options(s).

By Paul D. Tomlingson

World class performance is the ultimate objective of many maintenance organizations. It marks the organization as a leader in its industry and sets it apart as the ultimate achiever. But what is a “world class” standard and how can it be achieved? Is there a yardstick by which an industrial organization can measure the relative performance of its maintenance activity and compare it with performance levels that universally identify “world class” status? Are there financial formulas, achievements, evaluation techniques or indices that provide this yardstick?

Jan 1, 2006

By Brain E. Gombos

The Allen Park Sanitary Sewer Overflow (SSO) Tunnel and relief sewer project, located in the city of Allen Park, MI, is a long-term corrective action designed to bring Sanitary District One?s sanitary system into compliance with its 2005 consent order and service contract with the Detroit Water and Sewerage Department (DWSD). The $20 million project is intended to reduce Allen Park?s wet weather discharges to DWSD, reduce bypass pumping to the Ecorse Creek and limit the future risk of basement flooding by providing storage during wet weather events and eliminating hydraulic bottlenecks in the sanitary sewer system.

Jan 1, 2013

By Hendri Silaen, Charles Leatemia, Monang Napitupulu, John Kamarea, Arwin Edijanto, Riza Hasan, Avesiena Primadiansyah

The Deep Mill Level Zone (DMLZ) is the fourth panel cave mine in the East Ertsberg Skarn System, (EESS) which lies 500 m below the producing Deep Ore Zone (DOZ) mine and up to 1,700 m below the surface. The interaction between the DMLZ and DOZ is occurring as the DMLZ cave grows vertically toward the DOZ footprint. During this process, a rapid change in the global mine system stress is anticipated. This change is forecasted to drive substantial changes in the magnitude and direction of principal stresses in the DMLZ and DOZ mines at PT Freeport Indonesia. These changes will bring the mine system closer to the ultimate capacity of the excavations and can result in significant impacts to mine stability if not managed. Critical to this management, is understanding cave growth and locations, as well as how these changes will impact stress conditions.

Apr 26, 2022

Traditional loading systems have relied for many years on wheel and track loaders, power shovels, conveyor belts, and more recently, hydraulic front shovels. Each has certain advantages and disadvantages, depending largely on job conditions. Hydraulic backhoes were introduced several years ago as still another truck loading option. At first, buckets were small and backhoe use was limited to small earthmoving projects. Now, as backhoes are equipped with shorter sticks and bigger buckets to move larger quantities of earth. The "mass excavation" concept offers an excellent truck loading option in many situations. System Development and Advantages Mass excavation was first used successfully by contractors in site development and highway construction, loading both on-highway and off-highway trucks. Operators were able to take advantage of the excavator&apos;s ability to work in poor underfoot conditions and dig in tough materials. The technique was first applied to mining in the kaolin mines of central Georgia. Kaolin, one of the world&apos;s most abundant minerals, is used to make paper, paint, ceramics, rubber, plastics, and many other products. Since kaolin clay is slippery when wet, load-out with wheel loaders is very inefficient. Also, kaolin is found in very spotty deposits due to the weathering process involved in its formation. Consequently, draglines are also impractical for kaolin mining since they don&apos;t have the mobility to move between clay deposits. In response to these mining conditions and kaolin characteristics, mines began using excavators in mass excavation as early as 1976. They have developed a system that takes advantage of the excavator&apos;s •high penetration and breakout force in tight materials; •fast-cycling hydraulics; •ability to work in poor underfoot conditions; and •mobility in comparison to draglines. This loading method has been successful in Georgia&apos;s kaolin mines and now can be applied to other mining operations as well. The excavator&apos;s ability to carefully dig spotty ore deposits is an asset at uranium mining operations. Any mine can take ad-

Jan 3, 1982

By Verya Nasri, Heather M. Ivory, Kevin Clarke, Eric Cole

Development of the New York City Water Supply System The New York City water supply system serves a population of almost nine million, including residents in neighboring counties. Peak day demands on the system are currently running at about 70 m3/sec (about 1,600 mgd), which is significantly lower than in the 1990s. This trend of reducing demands over the past 25 years, which has been experienced in other major urban centers in the US, is mainly the result of conservation measures and increased cost consciousness on the part of water users. In the long term, demands are expected to once again follow an increasing trend with population growth, so that by the end of the planning horizon of these studies, the year2045, the peak day demand is expected to increase to about 85 m3/sec (2,000 mgd).The system is served largely by surface water sources, regulated by a system of 19reservoirs and three controlled lakes, which are connected to the City by a series of aqueducts, the longest of which is about 87 km (54 miles). The total storage in the sys-tem amounts to about 560 billion gallons, providing a system firm yield of about1300 mgd. The Delaware system provides about 50% of the City’s water, the Catskill about 40%, with the remaining 10% provided by the Croton. The system as a whole, which is shown on Figure 1, has been built up over the past 150 years, and is truly a marvel of engineering. The Catskill and Delaware aqueducts both discharge into Kensico Reservoir, which is a daily balancing reservoir located in Westchester County about 24 km(15 miles) north of the New York City line. The two aqueducts extend south for a further 14 miles to combine again at Hillview Reservoir in the Bronx, which provides hourly balancing storage for New York City system. The Croton aqueduct runs nearby as shown on Figure 2, but does not connect into the two reservoirs. From Hillview Reservoir the water is supplied to the city via three main distribution tunnels, known as City Tunnels 1, 2, and 3. City Tunnel No.1 was developed in the 1920s as part of the Catskill system. City Tunnel No. 2 followed in the 1960s as part of the Delaware system. City Tunnel No. 3 (CT3) is partly under construction and partly in operation, and is being implemented to permit the other distribution tunnels to be taken out of operation for the first time for inspection and maintenance. Presently CT3 is supplied with water from Hillview Reservoir via a large underground valve chamber, the Van

Jan 1, 2007

By Christopher J. Pritchard

Federal mine safety regulations prohibit the recirculation of face air In a gassy Class III metal-non metal mine. Tg Soda Ash, Inc, operates a 2.0M ton/yr (2.2 million st/yr) underground gassy Class III trona mine In southwest Wyoming, currently liberating approximately 5,700 m3/day (200,000 ft3/day) of methane. Because of the low methane emissions and the incombustibllity of trona, it was decided that Tg would be an excellent environment to test recirculation at a U. S. gassy mine, for possible future use by Industry. A research program was begun with the cooperation of MSHA, the U.S. Bureau of Mines, the University of Utah, and Tg. Initial testing centered around re-using shop air under¬ground. Next, conduct a continuous miner section recirculation test to see how the face contamination was affected by 305m (1,000 ft) and 610m (2,000 ft) recirculation distances. A district test followed to examine the effects of the critical parameter - respirable dust. Finally, a long-term program to study respirable trona dust transportation and deposition in mine airways to plan for district fan siting and effects on face air dust levels.

Jan 1, 1995

By William Lyman

GENERAL DESCRIPTION Shrinkage or shrinkage stoping refers to any mining method in which broken ore is temporarily retained in the stope to provide a working platform and/or to offer temporary support to the stope walls during active mining. Since ore "swells" when broken, it is necessary to shrink the muck pile a corresponding amount by draw¬ing some of the broken ore out as the stope is advanced-hence the name. Broken ore retained during stoping is drawn out after the stope has reached its limits. The stope may be left empty or may be filled with waste contemporaneous with, or subsequent to, the final draw. Traditionally the method implies conventional overhand stoping methods with miners working between the muck pile and the stope back, in a space which advances updip with mining and is maintained by balanc¬ing "swell" with "shrink." The shrinkage classification is also applicable to so-called "semishrinkage" methods in open pillar-supported stopes where broken ore is temporarily retained as a working platform but offers no wall support; and to various blasthole shrinkage methods which utilize broken ore temporarily retained in the stope for wall support, but which do not require miners to work from muck pile in the stope. The method is generally applied to steeply dipping veins of strong ore between strong walls. APPLICATION Geometry The geometry of a shrinkable vein is described in terms of dip, width, and regularity along dip. Overall strike and dip dimensions and irregularities along the strike generally impose no restrictions on the method. Dip is ideally 1.2 to 1.5 rad (70 to 90°). As dip falls below 1.2 rad (70°), the shrinkage draw begins to strongly favor the hanging wall side, thus leaving a poor working platform for conventional overhand work. This is particularly true in relatively wide stopes. The sup¬port afforded to the hanging wall also diminishes with decreasing dip, reaching nil as the dip approaches the repose angle of broken ore. Dips below 0.78 to 0.87 rad (45 to 50°) are not generally shrinkable except by open stope "seinishrinkage" methods. Minimum mining width is fixed by working space requirements in the stope-generally about 1 m. Shrink¬age in narrower veins requires that waste rock from one or both walls be broken with the ore and the attendant dilution accepted to achieve the minimum width. Nar¬row stopes are less suitable, encouraging hang-ups and bridging of broken ore, with the attendant problems of erratic draw and incomplete recovery of broken ore. Maximum practical width may be 3 m or less to over 30 m, depending upon the competency of the ore and its ability to stand unsupported across the stope back. This is a vital safety consideration in conventional over¬hand stopes, but is much less of a factor in blasthole shrinkage methods. Very wide veins and massive ore bodies have been mined by transverse vertical shrinkage panels separated by transverse vertical pillars which are either abandoned or recovered later by other methods. Regularity along the dip is a prerequisite of shrink¬age as there must be no serious obstruction to the flow of broken ore downward through the stope to the sill level. Gentle rolls along the dip are acceptable if the local footwall dip everywhere exceeds 0.78 to 0.87 rad (45 to 50°). Off-dip hanging wall and/or footwall splits can generally be mined selectively from a conventional shrink stope as they are encountered without ad¬versely affecting subsequent continuation of shrinkage mining updip on the main vein. Vertical offsets or major rolls along the dip which cannot be "smoothed over" generally require that a sublevel be established with new draw control development. Blasthole shrinkage methods are much less flexible (and thus less selective) in their ability to accommodate any of these irregularities. Ground Conditions The wall rock must be strong enough to stand with the minimal support afforded by the dynamic mass of broken ore in the stope. During active mining, local sloughing from the walls is restrained, but the broken ore affords little, if any, useful resistance to closure of the stope walls. Such squeezing, if present, may bind up the stope and cause the loss of much ore. Pillars left between and/or within stopes are effective in preventing closure but reduce overall recovery. Walls may be re¬inforced by bolting after each stope cut in conventional shrinkage but not in blasthole shrinkage. Ore in place must be strong enough to stand with no natural support across the stope width, although tem¬porary artificial support or reinforcement may be used locally in conventional stopes. Some spalling or sloughing is permissible in blasthole shrinkage as men are never present in the stope. Physical and/or mineralogical characteristics of the broken ore may impose restrictions on stope design and/or operational plan¬ning, and may even preclude the use of shrinkage al¬together. Examples include: ores which, when broken, are cohesive or which tend to pack or cement together under the influence of ground water, wall pressure, and/ or chemical reaction. Such conditions precipitate er¬ratic draw during mining and often result in difficult and/or incomplete final draw; pyritic ores which oxidize very rapidly in the stopes and may generate heat, imposing a fire hazard by spontaneous combustion; sulfide ores which oxidize sufficiently in the stopes to adversely affect mill recovery by flotation; and ores (es¬pecially those containing uranium minerals) which ex¬ude radon gas and thereby impose ventilation constraints on stope design. In most cases these problems can be minimized by limiting the size of stopes, by minimizing the duration of mining activity in each stope, and by promptly drawing each stope empty following comple¬tion of mining.

Jan 1, 1982

By Russel Delmar

Construction of Ontario Power Generation?s 12.8-m diameter, 10.4-km long water diversion tunnel in Niagara Falls, Ontario, is underway. The tunnel is being excavated by means of a 14.44-m diameter open gripper Tunnel Boring Machine. The tunnel lining is constructed as a two-pass system with the initial lining consisting of rock bolts, mesh, steel channels or ribs, and shotcrete, and the final lining consisting of pre-stressed cast-in-place concrete with a waterproof membrane. This paper describes salient design and construction aspects as follows: ? Regional and Site Geology ? Tunnel Layout and Design ? TBM Tunnelling and Cast-in-place Concrete Lining Construction

Jan 1, 2008

By S. R. Dindarloo

"Truck-shovel systems are the dominant material loading and haulage machinery in open pit mining operations. Optimal selection, sizing, and allocation of this capital intensive equipment is highly desirable for both achieving production targets and improving the mine’s economics. In this study, discrete event simulation technique was applied for modeling material transportation in Golegohar open pit iron ore mine in Iran. Direct observation of the material loading and haulage for duration of three months resulted in statistical modeling of the operation in the form of probability distribution functions. These probability models were used as the inputs of a stochastic simulation model constructed in GPSS/H simulation language. The model was validated through comparison with the actual operation and was used to test different what-if scenarios and sensitivity analysis of the current system. Some modifications in the number of trucks, shovels, and dispatching strategies were proposed. Simulation resulted in considerable improvements in the mine’s production and economics.INTRODUCTION There are three different mechanisms of discrete, continues, and hybrid material transportation in open mines, including but not limited to: i) in-pit crusher and conveyor belts (hybrid), ii) slurry piping (continues) and iii) truck-shovel system (discrete). Although, each mechanism has its advantages, the truck-shovel system is the dominant method of material loading and handling in open pit mines owing to its high production rate, excellent flexibility, relatively low operating and capital costs, and good maintainability. General objectives of optimal equipment selection include: i) meeting the long and short term requirements of production rates, ii) human & equipment safety, iii) environmental protection, and iv) economic operations (Figure 1).The two most important decision factors on a truck-shovel system selection and sizing are the equipment geometry and size. Geometry (equipment width, weight, turning radius, swing angle, etc.) is mostly controlled by mine’s designs and layouts, as well, as operational constraints. After selection of a favorable geometry based on the constraints; the next step is to select the equipment manufacturer. Once the geometry and manufacturers are selected; the next step is to decide on models (e.g. Bucket size) and required number of each unit"

Jan 1, 2015

By Nick Salisbury, Alan Howard, Scott Tunison

INTRODUCTION Geotechnical investigations for tunneling projects typically use vertical test borings and, to a lesser extent, angled borings. Horizontal drilling applications were often limited to specialty testing programs and to perform probe drilling from a tunnel boring machine (TBM) as part of production tunneling. The former often involved cumbersome setups with skid rigs and cribbing in a pilot tunnel or shaft. The latter usually was limited to generic probing for rock quality and ground water conditions with no sampling. Neither yielded continuous core at production rates suitable for design and construction planning. Conventional vertical investigations still make sense in many instances, particularly for tunnels planned beneath urban settings where truck access is convenient and where the tunnel horizon is not readily accessible via a portal or shaft. However, equipment is presently available to conduct horizontal geotechnicall investigations at reasonable production rates and cost. Advantages include continuously sampling the tunneling horizon and minimizing “access drilling” to get the zone of interest at the underground construction horizon. Horizontal boring lengths of up to approximately 800 m are considered feasible and those in the range of 150 to300 m are now routine. These capabilities make horizontal borings an attractive option to consider when looking at deep tunnels beneath ridges, especially where access is limited. Fewer drilling staging sites can also mean reduced environmental impacts when investigating tunnels proposed beneath sensitive lands such as public wilderness areas. Investment in early construction packages to develop portals for investigation access and, possibly in excavating test shafts for the same purpose should even be considered. Potential benefits in the form of achieving improved overall project costs using horizontal borings exist because of the significant additional data collected in the tunnel horizon. For this reason they should be considered during the routine cost-benefit analyses generally conducted for planning geotechnical investigations for tunneling. The authors also believe that collection of such data will reduce Geotechnical Baseline Report (GBR) conservatism and lead to lower potential for

Jan 1, 2005