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By George S. Barnett

The Equis and El Roble projects are located in the Cordillera Occidental of Colombia. The Equis vein deposits have sulfide ore reserves of 95,955 mt (105,770 st) grading 8.51 g/mt (0.248 oz/st) gold and 12.4 g/mt (0.36 oz/st) silver with 0.94 % copper, 1.10 % lead and 11.06 % zinc. Ore reserves at El Roble are estimated as 1.1 Mmt (1.21 million st) grading 4.89 % copper and 3.39 g/mt (0.099 oz/st) gold. Geologically possible reserves at Equis may total 1.6 Mmt (1.76 million st). El Roble has the potential to contain several tens of millions of metric tons. Exploration included diamond drilling, trenching, channel sampling, rock and soil geochemistry, geophysics and tunneling. Erratic high grade samples affect the weighting of channel and drill core samples. Secondary reconcentration of gold is also a factor. Sample maps and graphs show the distribution of values to be highly variable. Geostatistical treatment of some of these data gave important guidelines for parameters such as sampling densities, drill hole spacing and confidence limits.

Jan 1, 1985

By Dennis J. McCarthy

TAX TREATMENT OF EXPLORATION AND DEVELOPMENT EXPENSES Exploration Expenditures General Principles. Exploration expenditures are expenses incurred within the United States or the Outer Continental Shelf for the purpose of ascertaining the existence, location, extent or quality of any deposit of ore or other mineral. Section 617 of the Internal Revenue Code permits a taxpayer to elect to deduct these expenditures currently. In default of election, exploration expenditures must be capitalized and included in the basis of the mineral property for purposes of determining cost depletion. Entitlement to Deduction. In order to qualify for current deduction, exploration expenditures must be incurred prior to the commencement of the development stage, i.e., when mineral deposits are shown to exist in sufficient quantity and quality to reasonably justify commercial exploitation. Generally, expenditures for geological and geophysical investigations, reconnaissance, surveying, testpitting, trenching, drilling, driving of exploration tunnels and adits and similar types of work constitute exploration expenditures if incurred prior to the commencement of development. However, expenditures to further delineate the extent and location of an existing commercially marketable deposit are not exploration expenditures. Exploration expenditures may be incurred from within a producing mine to ascertain the existence of what appears to be a different ore deposit. Expenditures for property which is subject to an allowance for depreciation are not deductible under §617. However, depreciation properly allowable on assets used in exploration is an exploration expenditure. Election to Deduct Exploration Expenditures. The election to deduct exploration expenditures is made by deducting these expenditures in the return for the first taxable year for which such treatment is desired. The election applies to all properties of the taxpayer. Once the election is made all exploration expenditures, in the year of election and thereafter, must be deducted, unless the election is revoked with the prior consent of the Commissioner. The Tax Equity and Fiscal Responsibility Act of 1982 (TEFRA) made substantial changes in the treatment of exploration expenditures. For exploration expenses incurred after 1982 and before 1984 by a corporation, the amount currently deductible is limited to 85% of the expenditure. The remaining 15% must be capitalized and recovered over a five-year period. Although exploration costs capitalized under this provision are not part of the basis of the property for cost depletion purposes, investment credit may be claimed with respect to these expenditures. Further changes have been added by the Tax Reform Act of 1984 with respect to expenditures made after 1984. The Act reduces the amount of otherwise currently deductible costs from 85% to 80%. Additionally, it limits the availability of the investment credit for capitalized mineral exploration costs to deposits located in the United States. TEFRA also allows individuals a special method of recovery under which exploration expenditures may, at the election of the taxpayer, be amortized ratably over a 10-year period beginning with the year in which the costs are paid or incurred. If taxpayer does not use this method, the excess of the exploration expenditures actually deducted over the amount that would have been deducted if 10 year amortization had been elected is an item of tax preference subject to the alternative minimum tax.

Jan 1, 1985

By D. G. Wilder

D.G. Wilder I found the suggestion that the amount of displacement of a fault can be numerically related to the thickness of gouge or breccia to be both intuitively satisfying and intriguing. I have long agreed that there is some type of relationship between the amount of gouge and the amount of displacement of faults. I congratulate the author for developing a numerical relationship between them. However, I am concerned that the limits for applying this relationship be fully understood. An underlying assumption in this approach is that there is either a uniform thickness of gouge or breccia along a given fault or the thickness does not vary widely. Since it is not always possible to confirm this, the displacements derived by this method should be viewed with caution unless significant fault extent can be observed. At the Nevada Test Site, in drifts constructed in granite for test emplacement of spent nuclear reactor fuel, we found a fault with 0.3 to 0.4 m (12 to 16 in.) of clay gouge. Within a few meters of this location, the fault had no clay gouge, but rather consisted of a highly fractured zone with significantly altered rock and some slickensides. Based on Fig. 1, the 0.3 to 0.4 m (12 to 16 in.) thickness of gouge would indicate a displacement in excess of 30 m (98 ft). However, no gouge thickness would indicate essentially no displacement. Based on a quartz vein that terminated on the fault, and is not identified nearby, an estimated displacement of more than a few meters was made. This estimate is consistent with that obtained using the regression line proposed in the paper if the 0.3 to 0.4 m (12 to 16 in.) thickness for the gouge is used. However, using the regression curves with zero thickness would not yield results consistent with what was observed in the field. Therefore, it is important to recognize that the suggested procedure would properly yield a range of probable displacements. ? *Work performed under the auspices of the US Department of Energy by the Lawrence Livermore National Laboratory under Contract W-7405-Eng-48. Reply by E.C. Robertson It is certainly true that the t (thickness) of gg-bx (gouge and breccia) on a fault does vary along the fault. My observations have been that near the termination of a fault, the displacement d is small and the t is also small, whereas the maximum d and t will usually be found in the central part of the fault. The information on gg-bx and t of the fault found in granite in the NTS tunnel by Mr. Wilder could be interpreted somewhat differently than he does. He speaks of the fault changing within a few meters from 0.3 to 0.4 m (12 to 16 in.) of clay gg to "a highly fractured zone with significantly altered rock and some slickensides," but no gg. The highly fractured rock may be taken to be bx, rock not so finely ground as gg but still crushed by the fault movement, equivalent to the gg in my usage, and probably occupying about the same t. Mr. Wilder's estimates for the fault in the NTS tunnel for t of 0.3 to 0.4 m (12 to 16 in.) and for d of a quartz vein, in excess of "a few meters," would place the point on the low side of the central trend line in my Fig. 1, at the lower limit. There is, of course, a problem with determining d using displacement of only one planar surface. It would be greater or lesser depending on the rake of the movement. Finally, estimating the d of a fault from its t should be made with awareness of our present uncertainties, as pointed out by Mr. Wilder. Although the central trend line in my Fig. 1 has a ratio of d/t of 100, I have put the limiting ratios at 10 and 1000. Understanding of the values of the ratio will be improved only with collection of more data, for which the discussion of Mr. Wilder is much appreciated. ? G.C. Waterman E.C. Robertson's paper provides significant information to a geologist attempting to deduce fault offset by noting the products of structural dislocation. However, considerable mapping in underground and open-pit mines, and examination of structures produced in different geological settings, have convinced me that gouge and breccia thickness are controlled by geological conditions and fault movement. The following paragraphs suggest geological variables that control them. 1. Depth of Loading A near-surface fault resulting from tensional stress has more breccia/gouge than is produced by a similar stress at considerable depth. A deep-loaded compressional stress may produce a linear zone of schist, or structural dislocation may occur along an earlier formed belt of schist. Such "shear zones" are common in Canadian mines in precambrian rocks. In neither case can offset be directly deduced by an analysis of the minimal gouge/breccia in the shistose rocks. At greater depth, stress may be partially to wholly relieved by flowage. I vividly recall first noting the regional "Midas Thrust" in the Lark mine, Bingham Mining District, UT (where we called the structure the North Fault). My recorded notes, as I remember them, showed a narrow gouge streak separated the "Jordan" and "Commercial" limestone units from impure, muddy limestone beds of uncertain stratigraphic position. The visible structure did not indicate the great importance of this premineral fault

Jan 1, 1985

By Mark E. Emerson

INTRODUCTION A long time has passed since geologists encountered encouraging mineralization in their regional exploration program for base metals deposits. Detailed drilling during the following two field seasons served to delineate the configuration of the deposit. Various statistical techniques were employed to provide confidence levels regarding the extent of the reserves and to categorize them according to whether proved, indicated or inferred. At this juncture, it appears that the company has an orebody. Time elapsed? - three years; money spent? - $2.5 million. Much remains to be done before the orebody becomes a mine. The company urgently needs to develop new orebodies to replace its existing mines which are being depleted. It therefore has expended sizable amounts in exploration, the life blood of the company. A staff of specialists is maintained to evaluate new projects and to conduct feasibility studies. The scale of new mining projects today is beyond the resources of most mining companies. Bank financing, therefore, is often vital if the company is to survive. The following scenario describes how a company goes about evaluating mine projects and what is needed to meet lender's criteria in securing the needed financing. MINE PLANNING Now that the company engineers have taken over the project, their concern is no longer with finding ore but in developing it. An exploration shaft has been put down and a tunnel driven into the orebody to verify, on a large scale, what has been seen so far only in pencil- like drill cores. A large tonnage bulk sample is mined and shipped off to an independent metallurgical process laboratory for beneficiation testing. The budget is expended to cover engineering studies. Preliminary capical and operating costs are developed in- house and profitability studies are made based or1 assumptions concerning the markets for the main product and the indicated gold, silver, and other by-products. Time elapsed? - four years; cost to date? - $3.5-$4.5 million. The staff prepares its own pre-feasibility study. After review, the Executive Committee authorizes management to proceed with full-scale studies. The task is to make preliminary designs and to develop definitive capital and operating cost estimates. Specialized consultants will be engaged to study slope stability for the open pit and rock strength if underground recovery proves more feasible. If it is near the sea, an oceanographic report might be commissioned to see if the tailings can be disposed of off shore. Environmental impact studies are commissioned from a firm familiar with government requirements and regulations in this important area. The company subscribes for an expensive multi-client study of the supply- demand-price outlook for the product to support its own long-range price forecast. Now all the numbers are available to make a definitive appraisal of the value of the project. Cash flow analyses and sensitivity studies follow. Reams of computer printouts are generated. Time elapsed: five years. Amount spent? - up to $9 million, depending on the size, complexity and attractiveness of the project. Marginal projects always take longer. Finally, an engineering and construction (E&C) company is given the task of integrating and compiling all the data into a feasibility study. The results of dozens of reports, memoranda, computer printouts, and charts are boiled down to one final summary report of a few hundred pages together with conclusions and recommendations. After thoughtful review at a board meeting, the directors authorize management to proceed with detailed engineering work - provided that

Jan 1, 1985

By D. E. Spiller

Introduction Precious metals recovery from ore can generally be accomplished using gravity concentration, flotation, and/or hydrometallurgical (leaching) techniques. The objective of this paper is to show why gravity concentration can be an important part of recovery systems that employ leaching as the primary unit operation. A brief discussion of modem gravity concentration equipment is also presented. Discussion Gravity concentration of ores has generated increasing interest in recent years. Reasons for this interest include: • Gravity concentration is environmentally attractive. There is little or no use of reagents. Hence, it is relatively nonpolluting. • The cost of cyanide has continued to increase. Therefore, cost savings may be realized whenever leaching feed tonnage can be reduced by preconcentration. • Compared to flotation and leaching, gravity equipment costs are low per processed ton. Field installation costs for gravity circuits usually are less because many' units are supplied as self-contained modules. Also, the cost required to supply services, particularly power, to a gravity plant site are also less. In situations where preconcentration at coarse particle size is applicable, significant grinding equipment savings may be possible. • Gravity circuit operating costs are also relatively low compared to typical flotation and leaching circuits. Reagents, power, maintenance, and manpower savings in a well-engineered gravity plant may be realized. Again, if grinding is reduced, significant power and steel (media and liners) savings are possible. •In recent years, more efficient gravity concentrating devices have been developed. Benefits to Precious Metal Leaching Gravity beneficiation can complement precious metal leaching in two ways. First, the recovery of coarse liberated values before leaching may reduce leach time requirements and may reduce reagent consumption. Second, gravity preconcentration can reduce the size of a leach plant by decreasing the quantity of material to be leached. Coarse gold and silver have been shown to leach rather slowly. Kameda (1949) and Habashi (1967) have investigated the kinetics of cyanide leaching systems. They concur that in a heterogeneous reaction, the rate of gold and silver dissolution is directly proportional to the surface area. Thus, the instantaneous rate of dissolution for spherical 0.37 mm (400 mesh) gold is theoretically -25 times faster than for the same amount of gold at .841 mm (20 mesh), based on data from Fuerstenau, Chander, and Abouzeid (1979). Conversely, coarse liberated, +.841 mm (20 mesh), gold is more readily recovered by gravity concentration than is fine, -.037 mm (400 mesh) gold. Therefore, it is apparent that the two recovery systems complement one another. Figure 1 data demonstrates the potential synergism. A sample of - 3.327 mm (6 mesh) Nevada gold-bearing ore was cyanide leached using conventional bottle-roll test procedures. Gold extraction was determined as a function of leaching time. A second sample split from the same leach feed material was hand jigged to remove a coarse heavy mineral fraction, including virtually all of the +.210 mm (65 mesh) liberated free gold. This second sample, with the coarse gold and heavy minerals removed, was subjected to an identical cyanide leach procedure. Figure 1 presents the resulting comparative extraction data. Note that the percent gold extraction for the sample containing no +.210 mm (65 mesh) free gold includes the coarse gold recovered by gravity. The data show that the sample containing coarse gold required about 72 hours of leaching time to achieve 80% extraction. This compared to about 22 hours of leaching time for 80% gold recovery from the sample that contained only -.210 mm (65 mesh) free gold. Thus, there was a 69% reduction in leaching time. The improved extraction data is not wholly attributable to coarse gold removal, but rather it was the combination of gold removal and rejection of other heavy mineral cyanide consumers or leach retardants. Further investigation was not warranted at this time. Preconcentration is the second manner in which leaching systems can benefit from gravity concentration. The premise is that preconcentration can reduce the quantity of leach feed, which, in turn, may reduce leaching costs. Figure 2 presents preliminary data developed by CSMRI for US Minerals Exploration (USMX). Centennial Exploration Inc., in agreement with USMX, is proceeding with evaluations to determine the suitability of various processing schemes for recovery of gold values from the Montana Tunnels property. The data shows how the ore can be preconcentrated by gravity techniques to result in a reduced feed tonnage to secondary extraction techniques, presumably flotation or cyanide leaching. Testing has shown that Reichert cones, followed by treating the cone concentrate on spirals, can deliver about 88% gold recovery in about 13% weight, that is, 87% weight rejection. Consequently, fine grinding and reagent costs are attributable to only 13% of the plant feed rate. Cost data is not yet available, but the potential exists for significant cost savings.

Jan 1, 1985

By John Cogan

INTRODUCTION Lake tapping is a method of blasting an intake into a body of water from below the natural water surface without first lowering that surface or installing a protective cofferdam around the tap hole. Lake taps are done by first excavating a tunnel almost to the water/rock contact and then blasting out the final protective rock plug at one time to allow water to suddenly inflow into the tunnel from the lake. This procedure can be done "dry" or "wet." In the dry method the tunnel is empty of water before the final plug blast while in the wet method it is partially filled with water. This paper describes a wet method lake tap as performed at Alaska Power Authority's Tyee Lake Hydroeletric Project in southeast Alaska. The Tyee Lake project is located approximately 40 miles southeast of Wrangell, Alaska, in a glacially formed valley called the Bradfield Canal. The project consists of a powerhouse located at sea level on the canal, a system of tunnels and shafts connecting the powerhouse to Tyee Lake at elevation 373 m (1225 ft), and an 129 km (80-mile) electrical transmission line leading from the powerhouse to Wrangell and then Petersburg, Alaska. Tyee Lake itself is a natural lake gouged out of the rock by glacial ice during the last ice age. The objective of the project was to tap into the lake and use the water as a source for power production. The techniques used for lake taps have been developed in Norway where conditions are favorable for their use in hydroelectric proj¬ects. Over 100 have been performed. In the United States the Tyee

Jan 1, 1984

By Stefan S. Kapusniak, Nichol Riggott, Stephen F. Smith

Stability investigations have been carried out in a number of major coal mining tunnels in the U.K. using various instrumentatcon techniques. The paper describes these instrumentation methods and presents results from studies in several tunnels. Both long and short term stability aspects are discussed. Prediction of tunnel stability is examined in relation to tunnel support requirements. Computer simulation techniques are used along with actual underground measurements to identify tunnel design parameters for influencing choice of support and prediction of long term stability.

Jan 1, 1984

By David Branch, Martin J. White, Faramarz P. Hassani

The work presented in this paper is part of an overall research programme into the problem of stability of main access roadways driven in Coal Measure strata. Data about the post-failure behaviour of rocks are generally of a qualitative rather than quantitative nature. This investigation represents the initial part of a long-term research project aimed at quantifying the post-failure properties of Coal Measure rocks and to investigate their effects on tunnel stability. The results of tests conducted to evaluate the post-failure strength characteristics of Coal Measure rocks under uniaxial and triaxial stress conditions are presented. Empirical failure models are proposed to define the triaxial intact and residual failure properties of these rock types. The application of such data to closure prediction techniques of underground excavations are outlined and discussed.

Jan 1, 1984

By T. Krauthammer

A combined numerical approach is proposed for the analysis of rock-liner interact ion in tunnels under the effects of severe dynamic loads. The method is based on employing finite element formulations which describe the linear behavior and incorporating it into a finite difference code which can describe the behavior of a jointed rock medium. This method can be applied for the analysis of tunnel systems in rock, and for studying the requirements for adequate support of the tunnel.

Jan 1, 1984

By Michael B. Gross

A numerical study of the vulnerability of unlined tunnels in jointed rock has been performed with the STEALTH computer program, using the CAVS model for jointed rock. CAVS, for Cracking And Void Strain, is a constitutive model that can simulate the initiation/propagation of fractures and the slip along fracture planes in a jointed rock mass. The CAVS model is unique because it represents a spatially oriented approach to rock failure, rather than the traditional invariant formulation, and because the CAVS submodels for slip friction and slip-induced dilatancy are used to represent the weakening influence of joints, rather than applying ad hoc modifications to the elastic moduli of the host rock. The CAVS model, as embedded in the STEALTH computer program, has performed numerical simulations of tunnel collapse at high overstress levels. These calculations show that the ultimate strength of the tunnel depends strongly on the joint properties, and that the results of the calculations are in qualitative agreement with the observed data for collapse of a tunnel in tuff.

Jan 1, 1984

By John Kafritsas

INTRODUCTION Deformations and failure of rock masses in elopes, foundations or tunnels can be caused by water pressures. On the other hand, deformations of a rock mass affect the flow of water (and hence water pressures) by changing joint apertures. Including the coupling between flow and deformation effects is therefore important both in assessing the stability and deformation of a rock mass and in determining the amount of seepage taking place through it. The model that will be presented hare treats the coupling of the flow and deformation behavior of a rock mass. Deformations are modeled using the Rigid Block Method (Distinct Element Method) (Cundall, 19741, that is by assuming that the rock mass is composed of rigid blocks with deformable contact points. Flow is modeled by recognizing its discontinuous features, as occuring in a system of interconnected conduits formed by the joints. The program can also deal with problems involving free-surface flow. Example problems involving representations of rock masses are presented. Coupled flow modeling of interconnected conduits should be contrasted to equivalent-continuum models on the one hand and to pipe network models on the other hand. Continuous models substitute an anisotropically permeable continuum for the discrete flow that occurs in rock joints. Pipe or parallel plates represent the other extreme of existing models; they account for the discontinuous character of the flow problem, but cannot model rock mass deformation and failure. The coupled rigid block-flow model can eliminate most of these restrictions. The coupling between water flow and deformation is most pronounced in rock masses which, because of the arrangement of their

Jan 1, 1984

By R. G. Burdick, W. R. Peters

Shallow underground voids resulting from early coal mining and other resource recovery activities over the past several decades are now being recognized as a significant cause of ground subsidence problems. Uncertain knowledge of abondoned coal mines also imposes potential hazards in current underground mining operations, since water inundation or the release of methane gas is a serious hazard when mine excavation operations break into an abandoned mine. Requirements for detecting and mapping subversive manmade tunnels in previous work has led to the development of a surface-operated automatic earth resistivity survey system and associated computer data reduction process for military applications. Field demonstration tests of this system have resulted in successful detection of tunnels at depth-to-diameter ratios greater than 10 to one. Under the sponsorship of the Bureau of Mines, a similar high-resolution earth resistivity system was designed and constructed for use as a potential means of detecting abandoned coal mine workings. The hardware and software aspects of third new system are described together with applications of the method for surveying and mapping abandoned coal mine workings. Field tests are presented to illustrate the detection of both air-filled and water-filled mine workings.

Jan 1, 1984

By Simon T. Johnson

Potential applications for boom tunnelling machines requires reliable assessment of in-situ performance. The majority of existing classification schemes concentrate on the prediction of excavation rate, neglecting the important influence of tool consumption rates. A statistical approach is taken to the analysis of the performance data, which was obtained first-hand through in-situ monitoring, and laboratory testing of representative rock samples. A statistical discrimination technique is used to produce a scheme for the prediction of performance groups, each of which covers a limited range of cutting rate and/or tool consumption rate.

Jan 1, 1984

By Randall J. Essex

A massive mudslide blocked a river in Utah, impounding a 61 m (200 ft) deep lake. Under emergency conditions, a system was designed and constructed to: (1) drain the impounded lake; and (2) provide a permanent diversion for the river around the mudslide. The system consisted of a low level tunnel, a connecting vertical shaft daylighting close to the lake shore, control valves mounted within the tunnel, and an inlet channel between the lake and the shaft. This paper focuses on the problems and concerns surrounding the inlet channel excavation and the stabilization of the remaining vertical rock walls.

Jan 1, 1984

By Roger M. Zimmerman

The G-Tunnel heated block experiment is being conducted on the Nevada Test Site (NTS) as part of the Nevada Nuclear Waste Storage Investigations project (NNWSI). The purpose of the ambient temperature testing ph se is to evaluate rock-mass mechanical properties of a block (- 8 m3) under biaxial stress changes up to 7.5 MPa above an initialization in situ value of 3.1 MPa. Results indicate that the modulus of deformation ranges from 9.7 to 17.0 GPa and Poisson's ratio ranges from 0.21 to 0.33. In general, the higher values of the modulus and Poisson' s ratio were influenced by fracture propagations parallel to the compressive stress field. Other measurements indicated that cross-hole compression (p) wave velocities and single fracture permeability values were relatively insensitive to stress changes above the in situ value.

Jan 1, 1984

By C. M. St. John

INTRODUCTION The stress redistribution accompanying the excavation of a deep tunnel may induce failure of the rock. The role of the support system is to control the extent of the failed region and hence to limit closure of the tunnel within an allowable amount. In such circumstances, the design of the support system is generally based on the Ground Reaction Curve (GRC). The GRC is the relationship between the support pressure and the induced displacement at the tunnel wall for a circular tunnel under hydrostatic loading. In the calculation of the GRC (see Brown et al. (1983) for an exhaustive list of reference), the rock is generally treated as an elastoplastic material characterized by cohesive-frictional yield strength and by dilatancy during plastic deformation. A recent analytical development (Detournay, 1983) has allowed the generalization of the GRC concept for some cases of non-hydrostatic in-situ stress fields (Detournay and Fairhurst, 1982). As an application of that development, this paper describes design charts that provide powerful means of performing parametric analysis for tunnels subject to non-hydrostatic loading. BACKGROUND The in-situ stress in a plane perpendicular to the tunnel axis can be decomposed into mean and deviatoric components, Po and So (see Fig. 1). Since it is assumed that the rock is characterized by a Mohr-Coulomb yield strength, the deviatoric invariant So must be less than the limiting value Sl0 given by

Jan 1, 1984

By M. G. Schuerger, D. W. Park

An improved gage for monitoring fractures in underground openings was successfully developed using the reflection photoelastic technique. The gage may be used to detect crack initiation in mine openings or tunnels so reme¬dial action may be taken to prevent further deterioration. Fringes concentrated around the cracks were spread by using rubber backing. Crack growth as little as 76µm (0.003 in.) up to as much as a 2.54-mm (0.1-in.) width were detect¬ed in this study. Variation of gage materials will provide alternative ranges. The feasibility of the method was tested and calibration was made in a laboratory study. Field study successfully demonstrated the capability of applications in detecting mine roof crack development.

Jan 1, 1984

By Roger C. Ilsley

The $2.1 billion Milwaukee Water Pollution Abatement Program (MWAP) requires the construction of several miles of underground openings at depths up to 350 feet for the conveyance and storage of sewage and storm water overflows. The system will include approximately 27km (17 miles) of deep tunnels ranging from 3m (10 f t.) to 9m (30 f t.) in diameter, 26 shafts and a pumping station cavern. The facilities will be located primarily in Silurian Age dolomites of the Niagara Group. The rock is characterized as competent with vertical to sub-vertical orthogonal jointing and predominantly thin shale bedding planes. The site investigation was carried out incrementally, commencing on an area wide basis, becoming site specific as facility plans were developed. To achieve cost savings through uniformity and continuity, a computerized data bank was established. Software was developed during the initial stages of the program for the production of borehole logs, structure contour maps, geologic profiles and summaries of data for statistical evaluations. This paper discusses in general terms the collection, compilation and presentation of data necessary for the characterization of the rock mass.

Jan 1, 1984

By Z. T. Bieniawski

Tunneling in coal mines includes driving the main haulageways or roadways which serve as the development entries. Although constituting the lifelines of these mines, little attention has been paid to development entries in the United States in terms of systematic pre-construction planning and design. Coal mine tunnels, as the terminology is used in Europe and elsewhere, are understood in the United States to be the main development entries, either in room-and-pillar mining or in longwall mining. These development entries are not tunnels in the traditional sense since they usually are of a multi-entry type involving entries and cross-cuts, and their intersections. Thus, the design of coal mine tunnels is intertwined with the design of coal pillars. This paper summarizes the results of a three-year research project aimed at improving the design of development entries in U.S. coal mining. A design approach has been developed for this purpose and design procedures for roof support in main entries and for pillar sizing are proposed.

Jan 1, 1984

By H. H. Einstein

Swelling ground is one of the most feared problems in tunneling. Although usually not causing sudden failures during construction, the very large and longlasting deformations give contractors and designers the impression of facing an inevitable phenomenon against which there is no technical defense. Examples are concrete liners up to 1 m (3 ft) thickness that have been cracked or crushed, tunnel inverts that are heaving at 0.1 m (0.3 ft) per week during construction and continue to heave during decades. This paper summarizes essential components of design construction procedures where tunnels in swelling rock can be built to perform satisfactorily. It is important to point out that design and construction cannot and should not be separated. The way the tunnel will be built has to be an integral part of the design. Swelling can be defined as time dependent volume increase and is distinguished from squeezing, another time dependent phenomenon occurring in tunneling. Squeezing involves exceeding the shearing resistance while swelling usually does not. Swelling of rock masses can occur as swelling of the intact rock, swelling of wall rock, of gouge and of filler in the discontinuities of a rock mass, and as a combination of the above. Swelling mechanisms have been discussed elsewhere (ISRM 1983).

Jan 1, 1984