ABSTRACT: Fracturing in rocks is of major interest to the mining industry since, 1. the safety of mines depends, to a large degree, on the extent of fracturing, and 2. ore mineralization is found in the larger fracture planes. This paper is concerned with the safety aspect and presents the magnetic fabric method for determining in-situ values of the conjugate shear angle, a, and the angle of internal friction, F, which are important parameters for predicting fracture failure. The field example is a magnetic fabric study, involving rock samples from the surface and below the ground, in an area of southern Austria where rock bursting is common in mines. Samples taken from the mines for standard laboratory experiments employing a triaxial stress system show rock failure along axial extension fractures. The rocks, mainly dolomites, are too brittle under these conditions to fail in shear such that values of aand F cannot be determined experimentally. Using the magnetic fabric method in and around the mines, an in-situ value of a = 33 °and F = 24 ° was obtained. Furthermore, the degree of deformation within different rock types and directions of movements along shear planes were also determined.
1.1 Background and objectives
The close relationships between the strain- and the magnetic susceptibility anisotropy (MSA)-ellipsoids have been reported for rocks from many differing lithologies and origins. The two fabric ellipsoids reflect each other in their principal axes orientations and their forms; oblate (prolate) strain ellipsoids have co-axial, oblate (prolate), MSA- ellipsoids. Moreover, the data show direct correlations between the axial ratios of the strain and MSA-ellipsoids within a given lithology (see bibliography of Rathore,1988). In light of these close relation- ships, the speed, the ease and the reliability of the magnetic fabric method, it should be possible to employ the technique to provide additional information about certain tectonic or structural problems which cannot be fully resolved due to the lack of adequate geological information, specifically in regions where strains cannot be determined. This paper reports a magnetic fabric study from the Bleiberg region of Carinthia, Austria, where the main problem is to determine the extent to which the rock is strained. This problem is of importance because some parts of the lead-zinc mines in Bleiberg are very prone to rock bursts, due mainly to fracturing of bulky dolomite and limestone under stress. It is understood that the MSA method cannot predict actual stress magnitudes; however, it is be shown that MSA analyses can lead to the determination of the angle of internal friction, which is an important parameter in predicting rock failure.
1.2 Geology of the Bleiberg area
The lead-zinc mining area of Bleiberg is a part of the Eastern Gailtal Alps. The high valley of Bleiberg lies between the Drau and the Gall Valleys, which mark two of the most important faults within the Eastern Alps.
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.
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.
Mine gases cause major hazards to underground coal mining in many parts of the world. Among them, instantaneous outbursts of coal and gas as well as gas explosions are the most severe. During the formation of coal from vegetation, water, carbon dioxide and methane are produced in varying quantity. At the later stages of coalification, methane is produced in larger quantities and trapped within the coal.
It is commonly accepted that most of the amount of gases up till 80% - 90% of total quantity are deposited in coal or rocks as adsorbed ones, while the remaining 10% - 15% of the gas is recognized as the free gas. In theory, up to 200 m3 (Roberts 1983) or even up to 465 m3 (Kozlowski 1986) could be formed from 1 tonne of coal during coalification. The sorptive capacity depends on the internal area which in turnes is related to rank, to the nature of gas and to the pressure (Jackson 1984).
Gas hazard is a serious problem in Polish mining. Out of 70 coal mines operating in that country 44 are gasous mines giving 53% of total national production of coal. Accordingly, proper knowledge of gas content and its space structure in the deposit is of greate importance.
Very little has been done till now on study of space structure of gas content in deposits. Classifications of seams bases mainly on gas content hand produced maps with the use of simple algorithms of interpolation between data points (Kozlowski 1986). It has also been recognized as not creating truly reproducible results, i.e. results which are free from personal bias. Accordingly, to increase safety, high, single, occasional readings of gas content are considered decisive in the process of classification of gas hazard areas. Since the gas content data are very scattered, statistical methods to study the data should be employed. Among many of such methods (Agterberg 1974) geostatistics originating from the Matheron's concept of regionalized variable (Matheron 1903) has been successfully applied in many engineering areas (Journel and Huijbregts 1978). Structural analysis is one of the most important steps of geostatistical procedure.
2 GAS CONTENT AS A REGIONALIZED VARIABLE
There are many methods employed in determination of gas content in the deposit (Kozlowski 1986, Lama and Bartosiewicz 1983). The common future of the methods is their local character. Resulting from each method a numerical value of gas content is prescribed to a point in the deposit. The nature of gas distribution in the deposit as seen in Fig. 1 is typical to many geological values, for example : mineral grade, thickness of the seam, density etc. Regionalized variable theory as proposed by Matheron (Matheron 1963) is the basis for the ensuring geostatistical analysis.
Figure 1. Sample record of the gas content in the Lenin mine in E-W direction. (available in full paper)
A regionalized variable is any numerical function with a spatial distribution that varies from one place to another with apparent continuity whose changes cannot be represented by any workable function (Matheron 1970).
ABSTRACT: This paper introduces the concept of design theory and methodology as applied to rock mechanics for more innovative and efficient design. This is a "frontier" research area in rock mechanics because although design is a fundamental foundation to all engineering branches, very little attention has been paid to this aspect in mining. In discussing the theoretical premise of design, this paper identifies the principles which form the basis of design activity. Also, from a practical viewpoint, the findings from a series of interviews with mine design engineers are presented. These findings lead to the identification of needs arising from overlooked design methodology in practice.
In recent years, design theory and methodology have begun to emerge into a new discipline which supports all fields of engineering. However, concepts of design theory and methodology have neither been applied nor systematically studied in the field of mining engineering. This paper introduces these concepts as applied to rock mechanics for more innovative and efficient design, thus establishing a "frontier" research area. The importance of design research in the field of rock engineering is becoming increasingly evident. While it is true that there have been many impressive achievements in mining over the last several years, it is also true that innovations in mining have occurred at a much slower rate compared to other engineering disciplines. For example, it has been observed that the introduction of rock bolts in the 1940's represented the last major innovation in the area of strata control. Still today the layout of rock bolts is based primarily on empirical procedures. In summary, research into rock mechanics design is justified as follows:
ABSTRACT: Coal strength based on scaled uniaxial compressive strength from the laboratory and back-calculations from in-mine observations of pillar stability indicate that material strength scaling rules are open to debate. Analyses presented indicate that the application of scaling factors lower than 0.5 provide a better correlation with field observations of pillar stability.
Pillar sizing based on empirically derived formulae has been performed since the early 1900's. The current art involves the scaling of laboratory-derived coal strength to values representative of the in-situ coal mass. The various pillar-strength formulae, incorporating conventionally accepted scaling concepts, were applied to size pillars at an underground coal mine located in Pennsylvania. Even for presently stable pillars at the mine, the calculated safety factors were below those deemed adequate. Back-calculations based on observed pillar instabilities yielded an in-situ coal strength ranging from 730 psi (5.0 MPa) to 1073 psi (7.4 MPa) depending on the particular pillar-strength formula utilized. These values are approximately of the same magnitude as the coal strength determined in the laboratory [700 psi (4.8 MPa) to 1500 psi (10.3 MPa)]. As a consequence, the application of published scaling relationships to these laboratory data produced predicted in-situ coal strengths much lower than those that apparently exist based on the observed in-mine conditions. In this paper, a series of typical steps leading to determination of a "safe/stable" pillar size are applied. Discussion includes details relevant to the laboratory as well as field investigations, that generated concerns related to the use of common pillar formulae.
SPECIMEN PREPARATION AND TESTING
Although sample material for three coal sems (referred to as S-1, S-2 and S-3) was prepared and tested, it should be noted that the pillar sizing discussion will ultimately refer to only one of these seams, namely seam S-3. The samples for all three seams, as received, averaged one foot wide, two feet long, and one foot thick (across bedding). The sample material was obtained from recently mined area and delivered to the Penn State Rock Mechanics Laboratory wrapped in heavy plastic bags and burlap.
Table I, Results of Uniaxial Compressive Strength Tests on Coal(available in full paper)
Preparation of Specimens
The lump samples were kept moist and wrapped until specimen preparation, since it was expected that oxidation and progressive cracking from the drying of the coal material would deteriorate the structural integrity of the lump severely. Initially, specimen preparation efforts concentrated on obtaining cylindrical specimens using a standard laboratory coring techniques. It became apparent, however, that the specimen preparation techniques had to be altered for coal material from each seam. The standard coring technique provided useable test specimens for seam S-1 only. The material from seams S-2 and S-3 was very friable and fractured readily as coring was performed. Consequently, cubical test specimens were prepared from seam S-2 and S-3 material, exclusively. A horizontal band saw equipped with a tungsten-carbide blade was used to successfully prepare these cubical specimens. The ends of all specimens (cylindrical and cubes) were finished by hand with 200-grit sandpaper. Following preparation, specimens were kept moist in an environmental chamber (100% humidity).
ABSTRACT: A subsidence monitoring program over a longwall panel was established (1) to explore the impacts of dynamic subsidence on the ground surface and structures and (2) to correlate the movements between the structures and their corresponding points on the ground. The program consisted of a transverse line survey and a series of survey points around the structures on the ground surface and their corresponding survey points on the exterior walls of the structures. The vertical displacement of each point can be expressed as:
(available in full paper)
where Si is subsidence at i point; Sifis the final subsidence at i point; F is face location in terms of distance from the i point and SD is seam depth. The development of subsidence can be divided into five stages. Damage to the structures starts at the beginning of the second stage. The maximum damage developed in the middle of the third stage. The relationship of the vertical displacement between ground surface and structures can be expressed as:
(available in full paper)
where G is ground displacement, H is structures displacement, and a and b are constants. The relationship is linear. However, the strain is non-linear and more complicated because of the interaction of the ground surface and structure.
Traditionally nearly all subsidence research have been dealing with the final subsidence events, i.e. mechanisms, damages, and prediction of the final subsidence and paying little attention ix) the dynamic subsidence, i.e. the sequence of ground movement, subsidence development, and structural damages. Generally, both movements and damages during the dynamic period are more severe than the final stage. The objectives of this study are: (1) to explore the impacts of dynamic subsidence on the ground surface and structures, and (2) to correlate the movements between the structures and their corresponding points on the ground.
2 SITE DESCRIPTION
The study site is located in North-central West Virginia. Generally, the topographic relief in this area consists of gentle to steep mountains with meandering land between them. The mine is operating in [be Pittsburgh coal seam with an average mining height of 6.5 feet. The rock formations mainly consist of shale, sandstone, limestone, and coal. The thickness of the overburden is about 650 feet at the study site. The panel length is 4250 feet and width is 750 feet.
3 MONITORING PROGRAM
The monitoring program included (1) ground movement survey, (2) structure movement survey, and (3) crack mapping. For a period when the face was moving within i one seam depth from the transverse line, survey was conducted every other day. Figure l shows the locations of the survey line and stations. A transverse line (PW series of survey stations) was set up across the structures monitored and perpendicular to the face advancing direction. The spacing of the stations was about 50 feet, This line did not cover the whole panel width because access permits to some parts of the land were not available. A series of survey stations (P and W series) was installed on the ground surface around the structures i.e. at the corners and in the middle point of each wall. Crack survey consisted of recording all the cracks on, and damages to, the structures and ground surface.
ABSTRACT: Recent improvements are described in several techniques for ground control in underground mining. Promising methods for analysis and mitigation of rockbursts are discussed. Methods of assessing damage to underground excavations from repetitive seismic loading are reviewed. Analysis of mining-induced surface subsidence is discussed.
The purpose of this paper is to assess the current status of rock mechanics practice in mining and excavation engineering, and to identify those topics which need further attention to improve the reliability of particular aspects of engineering design in rock. The review is by no means comprehensive. It is intended instead to indicate some areas where substantial progress has been made recently, in both engineering principles and field practice, and to consider issues of topical significance in both mining practice and engineering construction in rock.
Mining practice involves maintenance or deliberate modification of the properties of the host rock mass. In this regard, research in techniques to preserve rock mass integrity has improved the technology of large-scale rock reinforcement and engineered backfills. Improved practices for design of grout curtains are developing for reduction of fissure permeability around shafts and similar facilities where groundwater flow must be restricted. For mine settings prone to induced seismicity and rockbursts, there has been considerable improvement in understanding of rockburst mechanics, with the prospect that control and mitigation measures may be developed.
There are several matters for which the state of theory and practice needs to improved to satisfy demands for a more reliable capacity to predict rock mass response to engineering activity. These include surface subsidence over longwall mines in areas with faulted cover and irregular surface topography, the dynamic performance of underground excavations subject to repetitive seismic loading and joint-controlled creep around excavations in hard rock.
1 GROUND CONTROL TECHNIQUES
Recent developments in ground control practice have been concerned with improvement of rock mass capacity to sustain induced loads and to maintain integrity while resisting displacements. Techniques which have improved substantially, in either operational function or design principles, include large-scale rock reinforcement, backfill design and grouting.
1.1 Rock reinforcement
Several recent well-executed investigations demonstrate the performance and benefit of cable reinforcement of stope boundaries. These include test stopes at the Cart Forks Mine (Pariseau et al., 1984), and the Mount Isa Mine, Australia (Greenelsh, 1985). The evaluation of several cable bolt reinforcement patterns in stopes at the Pyhasalmi Mine, Finland, is reported by Lappalainen and Antikainen (1987).
Field investigations of rock reinforcement show that the grouted steel tendons are effective when loads are mobilized in the reinforcement elements by inelastic strains in the host rock. This indicates that appropriate methods of design analysis for cable bolt reinforcement must provide for large strain in the constitutive model of the rock mass. A finite difference scheme simulating the interaction of deforming rock with grouted tendons has been described by Cundall and Board (1988), based on the conceptual model of reinforcement presented by St.John and Van Dillen (1983).
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.
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.
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.
The modulus of deformation of a rock mass is one of the critical parameters in the safe and cost effective design of underground structures. Traditional methods used to determine this property often involve laboratory testing of a group of core samples and "correcting" the laboratory results by using empirical factors related to joint spacing and other conditions. It is well known that core samples are representative of only a limited region in the formation and not necessarily of all the rock surrounding a tunnel. In addition, empirical correction factors have been developed from data having a substantial amount of scatter (Deere et. al., 1967). Other traditional methods such as plate bearing tests and the Goodman jack test provide only localized data and are subject to similar empirical correction factors (Bieniawski, 1978). As a result, traditional methods predict modulus values which are questionable and which err, in most cases, in a non-conservative way. Recently, under DEFENSE NUCLEAR AGENCY (DNA) sponsorship, UTD has developed a methodology for the measurement of the modulus of deformation through monitoring of convergence. This method recognizes that convergence of tunnel walls is the result of the elastic strain, and movement along joints and other anomalies, in thousands of cubic yards of rock surrounding the underground opening. Analytical solutions were derived which relate radial convergence to the elastic modulus (after the theory of elasticity [Timoshenko and Goodier, 1951]) of this large sample of rock, under conditions where induced stresses are not large enough to create a plastic zone around the tunnel. Since the method includes radial convergence which is a composite measurement of elastic strain of the rock, and other strain due to geologic anomalies, the term modulus of deformation is used (Bieniawski, 1978). The calculation of the modulus in this manner is analogous to calculating the modulus of elasticity of a complex composite material from strain and loading conditions. The advantages of using tunnel convergence over other methods is that this method is representative of a substantial "test specimen" complete with fractures, joint fillings, and other anomalies. It is, in fact, the modulus which exists where the rock formation interacts with the tunnel support system.
This paper presents the theoretical basis for the convergence method, and provides equations and charts that can be used to directly calculate the modulus of deformation from field measurements. Case histories are presented and emphasize the differences in magnitude between modulus values obtained from laboratory specimens, empirical methods, and the convergence method described in this paper. In one program UTD utilized the radial convergence measurement technique to obtain the elastic modulus of material at the Nevada Test Site. When the values obtained were used in design equations, they predicted tunnel behavior which agreed very well with experimental observations.
2 TUNNEL CONVERGENCE
Tunnel convergence is deformation caused by stress redistribution around the periphery of an opening during excavation. Considering the state of stress in an element on the boundary of an opening to be excavated, the state of stress prior to excavation is equal to the free field stress; i.e., the stress state of the element is equal to the undisturbed ground pressure.
ABSTRACT: A new approach to the direct inclusion of the effects of joints and geologic structure in 'rock mechanics analyses is applied to an elastic analysis of stability of a Lucky Friday undercut and fill longwall stope (LFUFL) using the finite element method (FEM). The approach is based on the concept of equivalent properties of a heterogeneous test volume, but does not depend on the assumption that a test volume is a representative volume element. Comparison with conventional FEM results show greater variability in fields of stress, strain, displacement, safety factor and energy densities. Stability requires a knowledge of rock mass strengths as well as elastic moduli.
This paper presents the results of a demonstration study of a new method for modeling the elastic properties of well-jointed rock masses. Joint is used here as a generic term and refers to the many structural discontinuities such as faults, bedding planes and so forth generally present in rock masses. Rock mass itself refers to field scale volumes of rock that are large relative to laboratory size test volumes. The latter are generally "intact" because they lack the discontinuities found in field. For this reason, rock properties determined from tests on laboratory sample volumes of intact rock and joints do not directly reflect the properties of the parent rock mass. As a consequence, the engineering analysis of stress and stability of rock masses remains highly subjective despite considerable advances in numerical analyses and computer hardware.
The demonstration study uses the UTAH2 finite element program and a method for estimating jointed rock mass properties recently developed at the University of Utah. The new technique, PM theory (Pariseau and Moon), is currently limited to modification of the elastic properties of rock masses. An outline of PM theory and several examples are given by Pariseau and Moon (1988); computational details for the elastic compliances can be found in the dissertation by Moon (1987). Extension to plasticity theory and strength estimation is described Pariseau (1988). Strength properties are used in the demonstration study but are not modified. The demonstration is two dimensional, although the overall approach is also applicable to three dimensional analyses.
The primary objective of the study is to demonstrate the practical feasibility of implementing the PM method in an analysis of stress about an excavation in a jointed rock mass. PM theory and numerical analyses of small scale two and three dimensional models of sample volumes of well-jointed rock masses show excellent agreement over a wide range of conditions. This is the first attempt to apply the PM method to an actual excavation in rock. The 5100 Level of the Lucky Friday Mine undercut and fill longwall stope (LFUFL) was selected for the study.
The approach to the study is a comparative one between two finite element analyses. The first is a conventional analysis based on average values of laboratory rock properties.