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Collaborating Authors
The 30th U.S. Symposium on Rock Mechanics (USRMS)
ABSTRACT ABSTRACT: The formation of tension cracks in slopes is well known. Analytical studies require the determination of a critical slip surface for any slope and, in doing so, it is necessary to consider an appropriate tension crack. A general yet simple approach is proposed for identifying the critical slope failure surfaces associated with critical tension cracks. The minimum factor of safety consequently reflects both safety against shear sliding and tensile failure. In general, the critical slope failure surface for a slope determined in this way differs from that which is obtained without optimizing the position and depth of the tension crack. In particular situations, the difference is significant. 1 INTRODUCTION After decades of research and investigation, a variety of optimization approaches have been developed for the analysis of slope stability. However, most of the approaches provide only a means for identifying the minimum factor of safety along simple surfaces. The common basis for such an assessment does not consider the likely position and depth of a tension crack. The effect of a tension crack on slope stability is usually taken into account based on an assumed or specified location and depth of the tension crack. The position and depth may be determined on the basis of stability charts( e.g. Hoek & Bray, 1977; Cousins, 1980), or empirical judgements. Analytical methods are also available for locating the tension cracks along assumed slip surfaces (Sancio & Goodman, 1979; Baker, 1981). For all these methods there are difficulties if a critical slip surface is to be searched without initial assumption as to the depth and location of the tension crack. Of course, if the position of a tension crack is known from its visible trace on the upper surface of a slope, the stability analysis of this slope should be primarily based on this exsiting information. However, when the critical tension crack position is unknown (i.e. there is no visible sign on the top surface or there are many tension fissures on the ground surface), it becomes necessary to use an optimization approach to locate the position and depth of the tension crack. In this paper it is proposed that the effects of tension cracks on slope stability be considered during the search for the critical failure surfaces. A simple and effective approach is presented for obtaining the critical surfaces associated with critical tension cracks. The development of the approach includes a modification of the Janbu generalized method of slices and the adaptation of a direct search technique for automating the iterative computations. In the proposed approach, no arbitrary restrictions are placed on the slope layering and the shape of failure surfaces, and the tensile strength of the rock or soil can also be taken into consideration. Moreover, although the methodology of the approach is mainly concerned with Janbu generalized method, it is applicable to any other commonly used method of slices for slope stability analysis.
ABSTRACT ABSTRACT: High-grade ore found in a shaft pillar at the Homestake Mine prompted a request for a Bureau of Mines study to determine if selected areas of the shaft pillar could be mined without jeopardizing the shaft. The design approach was to perform a finite-element study of the shaft pillar. Plan view and vertical sections were developed using two-dimensional, elastic, plane strain assumptions. Input into the model consisted of in situ stress measurements, geologic mapping, rock mass properties, and previous mining history. Results from the study indicated that ore-bearing sections of the shaft pillar could be removed within the displacement tolerances acceptable for the shaft. Homestake engineers, using the results of the finite-element study and considering other factors, have designed and are beginning to mine the shaft pillar. 1 INTRODUCTION Located in the northern Black Hills at Lead, South Dakota, the Homestake gold mine has been in operation since 1876. Current annual production is 1.8 million tons and development extends to more than 8,000 feet below the surface. Vertical crater retreat (VCR) mining accounts for approximately 60% of total underground production (Muir, 1988), with the remainder coming from mechanized cut-and-fill methods (Smith, 1987). Concerned with stope and pillar sizing for VCR mining in a large ore block at increasing depth, Homestake entered into a cooperative rock mechanics research program with the Spokane Research Center of the Bureau of Mines and the Department of Mining Engineering of the University of Utah. The extensive field investigation became part of the Bureau's research program for developing advanced concepts for mining deep ore bodies. The success of this study, which was conducted on a panel between the 6950 and 7100 levels, led to a second cooperative agreement between Homestake and the Bureau (see Pariseau, 1985). The Ross Shaft is one of two main shafts servicing the Homestake Mine. Approximately 50% of all ore is transported through the shaft, which also contains the main pump column and half of all the other service lines for the mine. During sinking operations in the early 1930's, a large ore body was intersected between the 3200 and 3800 levels. Subsequent development proved the existence of a large ore zone, and a shaft pillar was defined. The pillar, which extends for 61.9 m along the strike of the ore zone and is bounded by geologic contacts between the ore body and the Poorman and Ellison Formations, contains 8.8% of the mine's proven gold reserve. Production goals dictated mining the pillar while maintaining the integrity of the Ross Shaft (Corso, 1988). The goals of the cooperative program were to: (1) provide an instrumentation plan to monitor shaft and pillar loads and deformations, (2) provide input for calibrating computer models, (3) develop a reliable computer model to be used for evaluating shaft stability, and (4) provide guidelines for stope sizes and sequencing. 2 APPROACH Although a precise measurement of the amount of shaft deformation that can be tolerated has not been determined, it appears reasonable to suppose that as long as deformation is within the elastic range, displacements will be tolerable.
- North America > United States > South Dakota (0.26)
- North America > United States > Utah (0.24)
- Geology > Geological Subdiscipline > Geomechanics (1.00)
- Geology > Mineral > Native Element Mineral > Gold (0.54)
ABSTRACT ABSTRACT: Much of the present knowledge about caving behavior of rock masses has been obtained from empirical observations. Additional notions about caving have been developed through inferences derived from two-dimensional finite element analyses. These analyses have indicated that a combination of one low-angle set of fractures and one nearly vertical set of fractures is the optimum fracturing configuration for ease of cavability of an orebody. This paper presents the results of two- and three-dimensional distinct element analyses which draw different conclusions than those reported from finite element studies. The distinct element method is selected for analysis of cavability because this method treats the rock mass as an assemblage of rock blocks which may interact individually. The results of the analyses are compared to a documented case history which involved a groundfall of 80,000 tons of ore in a 160-foot high pillar. The mechanics associated with these results are explained in terms of simple static stability analysis of wedges. The propensity for orebody caving is primarily a function of the number of joint sets or potential "release" surfaces in the orebody. This mechanism is influenced by the bounding weak discontinuities and intact strength of rock material. 1 INTRODUCTION The cavability of orebodies is important to various mining methods in distinctly different ways. The block caving mining method relies on caving to extract massive ore economically, whereas other methods rely on the stability of the orebody and host rock to extract ore selectively. Cavability is a function of the geomechanical properties of the rock mass and the in-situ and mining-induced stresses. It has long been recognized that "the ability of a block to cave or fragment is a function of its strength in tension or shear and the value of applied forces" (Bucky, 1956). Numerous attempts have been made to develop classification systems for use in determining cavability. Much of the present knowledge about caving behavior has been obtained from empirical observations. For example, Mahtab and Dixon (1976) concluded from observations that the principal geomechanical features influencing cavability are in-situ stress field, rock strength, and the geometry and strengths of discontinuities in the rock mass. These authors also, by back-analysis of elastic calculations, postulated effective fracturing configurations. They concluded that "a combination of one low-angle (0° to 30° dip) set of fractures and another nearly vertical (75° to 90° dip) set of fractures is the most effective two- dimensional fracturing configuration for ease of cavability of an orebody. In an actual three-dimensional situation, one set of low angle fractures and two sets of nearly vertical fractures will be most effective in improving cavability." These observations, concerning favorable joint orientations, may be valid for environments lacking lateral confinement (i.e., as a result of boundary slots or boundary weakening). However, results of two- and three-dimensional distinct element analysis, presented herein, suggest that caving in confined environments requires that additional release surfaces be present. Release surfaces may be either preexisting or form as a result of high horizontal stresses.
ABSTRACT ABSTRACT: Stratabound copper ore bodies in thick Proterozoic metamorphosed sedimentary and volcanic rocks at the Tong Kuang Yu Mine in southern Shanxi Province of China were considered as possibly being amenable to block cave mining. Initial evaluation of caving characteristics was based upon old calyx drilling records, limited underground mining experience and geological mapping, a few diamond drill holes and measurements in surface subsidence features. Classification and mapping distribution of rock quality types aided in mine design, prediction of fragment sizing and explosive requirements, cave assist and ground support requirements. Comparison is made between early feasibility predictions and actual mine experience. 1 INTRODUCTION Evaluation of mineral deposits in developing nations often requires decisions regarding mining methods and rock behavior on the basis of limited data, limited drilling and testing and limited budget. Under these conditions it is particularly important to utilize all available geological and drilling information and establish correlations with probable rock behavior during mining. This was the situation at the Tong Kuang Yu Mine feasibility study. Mining of the Tong Kuang Yu deposit, located in mountainous country in the southern Shanxi Province in China (fig.l), began in 1960 following a 1955-56 calyx drilling Program. This early drilling consisted of 50,000m which discovered 100 separate lenticular concentrations of disseminated copper mineralization. Ore bodies #4 and #5 were the largest with low-grade reserves of approximately 130Mt above the 690m drainage level and plus 170Mt below the 690m level. Mining operations since 1960 have been concentrated above the 930m level using sub-level caving methods. Production was dropped to the 690m level and transported 5km to the concentrator. Figure 1: Location of Tong Kuang Yu Mine(available in full paper) Early in 1984 a delegation from the Sundt Corporation and Sunshine International made a presentation to the engineering staff of the TKY Mine and the China National Non-ferrous Import and Export Corporation regarding advantages of block caving as an underground mining method for certain types of large low-grade deposits. A feasibility study from TKY data led to engineering and construction phases for #5 ore body conducted under tight, budgetary restrictions. 2 GEOLOGY The ore bodies occur in thick metamorphosed sedimentary and volcanic rocks and are stratabound in units locally called "Metagranodiorite", which appears to be a metamorphosed acid tuff, and to a lesser extent in "Metadiorite" which appears to be a metamorphosed andesite tuff. The mineralized tuff units are interbedded with and overlain by quartzite and sericitic quartzite and the sequence is underlain by chlorite schist (fig.2). Figure 2: Plan and section of the geology of the TKY Mine area.(available in full paper) The structural history has been complex. The resultant is a northward plunging antiform structure (probably an overturned syncline) bounded on the east by a major north-northwest trending fault zone, the Tong Kuang Yu Fault. Isoclinal folding, development of schistosity and metamorphism of the units took place during Proterozoic time. Copper mineralization was essentially contemporaneous with deposition of the tuff members but suffered some redistribution during metamorphism. Unmetamorphosed and unmineralized diabase dikes were emplaced probably during Jurassic time.
- Geology > Mineral > Native Element Mineral > Copper (1.00)
- Geology > Geological Subdiscipline > Geomechanics (1.00)
- Geology > Geological Subdiscipline > Volcanology (0.95)
- Geology > Structural Geology > Tectonics > Compressional Tectonics > Fold and Thrust Belt (0.54)
ABSTRACT ABSTRACT: Based on the results of in-situ observations and physical model analysis, a computer simulation method, FEAEBP, has been developed to predict the behaviors of main roof breakage in longwall mining by considering it as a Kirchhoff plate on Winkler elastic foundation. This method is used to investigate the initiation, and development of the breaking process of main roof and its displacement field before and after its breakage. In this paper the simulation method is introduced, characteristics of the displacement field of the main roof is discussed and monitoring variation of the displacement field of the main roof in a Chinese longwall face is demonstrated. 1 INTRODUCTION It is well known that the main roof should be induced to break in longwall mining. Without it the main roof will cause roof weighting which may endanger the miners and mining operations. Normally, in China a longwall face is about 100-200 m wide and the thickness of the main roof is usually 2-8 m. Thus, the main roof in this case can be considered as an elastic plate, or a Kirchhoff plate, when it is separated from the upper strata or when the friction force between them is very small. On the other hand, the immediate roof and coal seam around the plate, being served as supporting abutments for the main roof, can be considered as an elastic medium ix) conform with the assumption for the Winkler elastic foundation. Based on these assumptions, the theoretical model of main roof can be simplified as a Kirchhoff plate on Winkler elastic foundation (Fig. 1), and a computer simulation method, FEAEBP, has been developed. This model is capable of simulating a moving face under different boundary conditions. FEAEBP has demonstrated that it can effectively simulate the breaking process of the main roof (1,2). Using this model the displacement fields of the main roof before and after its breakage were obtained, and some new concepts found which are introduced in this paper. 2 DISPLACEMENT FIELD OF MAIN ROOF BEFORE ITS FIRST BREAKAGE In the middle of a longwall face, the mode] can be simplified as a beam on Winkler's elastic foundation (Fig. 2(a)), the displacement, y, of the main roof over the seam is derived by the following equation (1) EIy() + ky = q and boundary conditions (2) EIy() = M x = 0 (3) EIy() = Q x = 0 (4) y = q/k x = 8 where M ---- Internal bending moment of the main roof at x = 0, Q ---- Internal shearing force of the main roof at x = 0, K ---- Composite elastic coefficient of the immediate roof and the seam. Solving for Eqs. 1 - 4, (available in full paper) where ß = 4vk/4EI --- calculation parameter, Fig. 2(b) is the displacement curve of main roof in Yunggong Mine by substituting into Eq. 5 the following parameters obtained through field observations and laboratory testings:
ABSTRACT ABSTRACT: A users friendly PC Computer Program has been developed to implement a systematic design of the powered support. The required input data are: (1) thickness, RQD, and compressive strength of the immediate roof; (2) mining height; and (3) type, thickness, and tensile strength of the main roof. Several important parameters for support selection under a given geological condition are generated through the implementation of the program, including type of support most favorable for the geologic condition, minimum inclination angle of the legs with respect to the horizontal axis, suitable setting load ratio of the front to rear leg for the 4-leg supports, optimum setting load, and support capacity. 1 INTRODUCTION Determining an optimum capacity of the support and selecting a proper type of support are two major problems in the selection of powered supports at longwall faces. Based on field instrumentation, finite element and analytical analyses, the methods of determining the rational load capacity of support and selection of suitable support type have been developed (Peng et al., 1987; Hsiung et al., 1988; Peng et al., 1989). The methods are simple yet reliable. In this paper, a users friendly PC computer program for implementing the systematic design of the powered support is introduced. 2 BASIC METHODS OF SUPPORT DESIGN 2.1 Major factors to be considered for support design Previous studies showed that the following seven factors must be considered in determining the support capacity and type: (a) thickness, (b) uniaxial compressive strength, and (c) RQD of the immediate roof, (d) ratio of the immediate roof thickness to mining height, and (e) type, (f) thickness, and (g) tensile strength of the main roof. Different combinations of those seven factors required different support capacity and type. 2.2 Selection of support type A method of selecting the suitable type of powered support at tong wall faces have been developed (Hsiung et al., 1988). The major steps for selection of support type are: (1). Determining the roof stability index In general, different roofs behaves differently. Different kinds of ground control problems will result from different roof conditions. Based on those seven factors and the weighting factors assigned to represent the relative importance of each variable, a roof stability index Q, was developed (Fig. 1). It ranges from 0.5 to 8. A roof with a large index value is a weak roof while a small index value means a strong roof. (2). Evaluation of roof characterization As mentioned above, different roofs behave differently. For example, the features of a weak roof are: the roof falls very easily in unsupported area between the canopy tip and faceline and tends to break into small sizes of rocks, which can easily enter into the working area and cause problems if a support with no shield protection is used. However, under a strong roof condition, roof falls in the unsupported roof area is no longer a problem since a strong rock has a higher tensile strength, but it tends to overhang behind the support into the gob.
- Materials > Metals & Mining (0.47)
- Energy > Oil & Gas > Upstream (0.35)
ABSTRACT 1. INTRODUCTION The excavation of a longwall panel causes a continuous stress redistribution in the surrounding strata for every face advancement. The understanding of the behavior of rock to high level of stresses is instrumental to the structural design of underground mines. In order to achieve a realistic analysis of the behavior of strata, the three-dimensional finite element method is frequently used. Also the model developed for the analysis has to be designed large enough to represent the influenced strata in the study area and a sufficient number of elements has to be used to provide detailed information on the stresses and displacements. Due to the size of the model and its numerical solution the CRAY X-MP/24 supercomputer has been used for this study. The mine site selected for the analysis is a coal mine located in the Black Warrior Basin in Alabama. The mine has two longwall sections at a depth of 2,000 feet with various dimensions of chain pillars, including yield pillars. Various research programs at this mine have been carried out by Park et. al. ( 1984 ). Therefore data were already available on the geology, and physical properties of rock and coal. In addition a rock testing program has been carried out to obtain rock mass classification parameters. The progressive failure technique has been utilized to simulate stress redistributions. The failure criterion after Hoek and Brown ( 1980 ) which is based on rock mass behavior was used for the iterative technique. Simulated stress redistributions are therefore very realistic. In order to automate the process involved in the application of progressive failure technique, a computer program has been developed. The combination of a large scale three-dimensional finite element model and the automation of progressive failure simulation provided a tool that can generate a large amount of information related to the behavior of rock and coal as a response to a longwall mining operation. 2. FINITE ELEMENT MODELING The three-dimensional finite element model covers an area of 1,160 feet * 1,050 feet which consists of three rows of pillars and two half longwall panels, with a half panel width of 310 feet, situated on both sides of the chain-pillar entry system ( Fig 1. ). The total height of the model is 2,130 feet accounting for an overburden above the coal seam of about 2,000 feet and incorporating two layers below the coal seam. The model is discretized vertically into 14 layers of strata groups. This gives a practical representation of the stratigraphic sequence above the mine. Each layer consists of 531 eight-noded hexahedral elements, resulting in 7434 elements in total. In the area of the panel face the finite element mesh has been refined to allow a detailed analysis of stresses in the pillar corners and at the longwall panel face. Boundary conditions for the model were imposed with single constraints in the x-direction for the two vertical y-z boundary planes and in the z-direction for the two vertical y-x boundary planes. The bottom x-z boundary plane was constrained in the y-direction.
- Geology > Rock Type > Sedimentary Rock > Organic-Rich Rock > Coal (1.00)
- Geology > Geological Subdiscipline > Geomechanics (1.00)
ABSTRACT ABSTRACT: This paper summarizes ground control studies of longwall panel entry systems in two Western U.S. coal mines. Presented are comparisons of in situ pressure change measurements and assessments of chain pillar and entry behavior under a wide range of cover. Results indicate that no one entry system design is universally applicable, and that site-specific factors need to be considered for panel entry design. Relationships between overburden depth, face advance, and magnitude and location of the forward abutment are discussed for western operations. In addition, analysis indicates that use of yield pillars may reduce or eliminate some stress-related problems that plague longwall panel entry design. 1 INTRODUCTION The objective of this research is improved understanding of longwall mine stability problems under various conditions. The ability to deal with structural stability problems in longwall mining and associated risks to mine personnel is highly dependent on understanding stress redistribution during different stages of mining. Many factors contribute to high-stress concentrations, such as mine design, mining sequence, and site stratigraphy and its reaction to variable loading during distinct stages of mining. As geological conditions, physical properties, and mining practices and face support capabilities may vary from mine to mine, it is difficult to determine a single generic stress pattern that would be applicable to all coal mines. This report briefly describes hydraulic borehole pressure cell measurements made in two Western U.S. longwall operations as part of previous and ongoing Bureau studies (1-5). This summary presents data from mines representing the wide ranges of depth in the Western United States. While field investigations usually include a wide variety of tests and in situ measurements, this paper utilizes pressure cell changes and concentrates on forward and side abutment pressures and entry behavior, with an emphasis on longwall-induced load transfer onto chain pillars. Pressure change data from longwall mines with two and three entries are presented from instrument sites at overburden depths ranging from 450 (137 m) to 2,000 ft (608 m). 2 MINE SITES DESCRIPTION This study summarizes data from two longwall coal mines typifying the wide range of Western U.S. mining conditions. Mine A, under relatively shallow cover, 450-750 ft (137-228 m), utilizes both two and three-entry longwall panel systems. The main roof in mine A is composed of variably thick sandstone units, and the immediate roof is sandy shale or siltstone. Approximately 1 ft (0.3m) of carbonaceous shale overlying up to 20 ft (6 m) of sandstone comprised the immediate floor. Mine B, under relatively deep cover, up to 2,500 ft (760 m), historically experienced coal bumps due to rapid changes in overburden depth and the presence of a strong, thick sandstone roof member. Longwall mining at this mine utilizes two-entry systems with yielding chain pillars. Immediate roof strata consists primarily of a regularly caving sandstone member. The floor consists mostly of sandstone, shale, and coal. In both mines the seam heights are comparable; other pertinent study site geometries and properties are shown in tables 1 and 2.
- Geology > Rock Type > Sedimentary Rock > Clastic Rock > Sandstone (1.00)
- Geology > Rock Type > Sedimentary Rock > Clastic Rock > Mudrock > Shale (0.95)
- Geology > Rock Type > Sedimentary Rock > Organic-Rich Rock > Coal (0.90)
ABSTRACT ABSTRACT: A method is proposed to predict the shear response of a dilatant rock joint under a wide variety of boundary conditions from the results of conventional direct shear tests. The validity and applicability of the method are discussed. 1 INTRODUCTION The response of a rock joint to shear loading in situ depends to a large extent on its surface properties as well as the boundary conditions that are applied across the joint surfaces. The range of boundary conditions can best be represented by assuming that the deformability of the rock mass surrounding the joint is modelled by a spring with stiffness K = dsn/dr where dsn and dv are the changes in joint normal stress and displacement, respectively. The applied stiffness varies between zero for a joint under constant normal stress (as in slope stability problems) and infinity if the rock mass is very stiff for which no change in joint normal deformation is allowed. The stiffness is constant if the change in joint normal stress is proportional to the change in joint normal displacement. The stiffness may also vary with the load history of the rock mass as it undergoes cycles of loading and unloading. The range of joint normal loading conditions in situ and the importance of properly modeling rock joint behavior have been emphasized by Goodman (1976), Heuze (1979), Leichnitz (1985), Goodman and Boyle (1985) among others. Lam and Johnston (1982) also recognized the importance of properly modeling the boundary conditions at the interface between concrete and rock when assessing the side resistance induced in a concrete pile in rough rock sockets as the pile is loaded vertically. Joint response under constant or variable normal stiffness boundary conditions has not received much attention in the literature. Test data are limited (Leichnitz, 1981,1985; Hutson, 1987; Lam and Johnston, 1982) since constant or variable normal stiffness shear tests on rock joints are difficult and require complicated computer controlled equipment. In a recent paper, a method was proposed by Saeb and Amadei (1988) to predict the response of a dilatant rock joint to shear loading under constant or variable normal stiffness boundary conditions. The method just requires knowledge of the behavior of the joint under constant normal stress, obtained from conventional direct shear tests. The purpose of this paper is to further discuss the validity and applicability of the method. At the outset, a brief description of the method is presented. Then, the method is verified using experimental data available in the literature. Finally, the applicability of the method to assess the effect of rock mass deformability on the stability of a block in the roof of an underground excavation is discussed. 2 JOINT BEHAVIOR UNDER CONSTANT OR VARIABLE NORMAL STIFFNESS 2.1 Presentation of the method Consider the "idealized" response of a rock joint to shear under constant normal stress boundary conditions represented by the solid lines in Figures la, lb and lc. These curves were proposed by Goodman and Boyle (1985) and are used in this paper with the sole purpose of introducing the method.
- North America > United States > Colorado (0.29)
- Europe (0.29)
ABSTRACT ABSTRACT: Seismic and radar tomography of the SCV underground test area in Stripa, Sweden have detected two major sets of fracture zones. Slickenside striations in these zones indicate that they have undergone shear deformation. Utilizing a database consisting of 3100 logged fractures from borehole core and 900 fractures mapped on the walls of underground drifts, the character of these fracture zones have been analyzed. This data along with numerical modelling suggest that the higher fluid conductivities in the zones compared with the surrounding rock may be due to a combination of higher fracture densities, and stresses in the zones that promote dilatancy in fractures with certain orientations. This would also result in anisotropic flow in the zones, with the primary flow direction in the direction of the dilatant fractures. 1 INTRODUCTION Stripa, Sweden is the site for a test area regarding the underground storage of nuclear waste in crystalline rock. This test area is located in underground openings in Stripa granite, a generally unfoliated but highly fractured rock (Carlsten, 1985). The most recent study area at Stripa, referred to as the Site Characterization and Validation (SCV) site, consists of a volume of granitic rock approximately 150 m by 150 m by 50 m at a mean depth of 360 m, referred to as the SCV block. The SCV block is being characterized using geologic, seismic, radar, and hydraulic data. Radar and seismic tomography of the area have detected two major sets of fracture zones. The first set strikes roughly north-south and dips steeply to the east, and the other set strikes northeast and dips 35-40 °to the southeast (0lsson et al., 1988). The approximate location of zones from these two sets in the SCV site are shown in Figure 1 (these zones are not in exactly the same location as given in Olsson et al., 1988, see below). Figure 1 also shows the location of five boreholes drilled through the SCV site and underground drifts that border it. Core logs from these boreholes show that the zones vary in width from 2 to 15 m, and have a higher fracture density than the surrounding ground (fracture density in the zones varies from 10 to 30 fractures/m, compared with 1 to 20 fractures/m in the ground surrounding the zones). Also, there is evidence that the hydraulic conductivity can be higher in the fracture zones than in the surrounding ground (0lsson et al., 1988) Figure 1 360 m level at Stripa showing the location of the SCV site and the fracture zones identified by radar and seismic tomography.(available in full paper) Lawrence Berkeley Laboratory (LBL) is developing a conceptual hydraulic model for the SCV block. The fracture zones shown in Figure 1 are considered to be the primary conduits for flow in this model. The location of some of the zones in Figure 1 have been modified from the location as given in Otsson et at., 1988, in order to better match with the results of hydraulic tests conducted in the SCV block.
- North America > United States (1.00)
- Europe > United Kingdom > England (0.28)
- Geology > Structural Geology > Tectonics > Plate Tectonics (1.00)
- Geology > Geological Subdiscipline > Geomechanics (1.00)
- Well Completion > Hydraulic Fracturing (1.00)
- Reservoir Description and Dynamics > Reservoir Characterization > Faults and fracture characterization (0.96)
- Reservoir Description and Dynamics > Unconventional and Complex Reservoirs > Naturally-fractured reservoirs (0.76)
- Reservoir Description and Dynamics > Reservoir Characterization > Reservoir geomechanics (0.70)