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Rock bolting is one of the most effective and economical means of rock mass reinforcement. Existing studies of rock bolt reinforcement are mostly focused on rock masses without flaw, with a single flaw, or with parallel flaws. However in rock masses, cracks or flaws usually exist in the form of cross-flaws. In order to understand the impact of cross-flaws on rock bolt reinforcement and to further explore the differences of bolt reinforcement between rock mass with cross-flaws and rock mass with a single flaw, reinforced analog specimens with cross-flaws and with a single flaw were tested under uniaxial compressive condition. The experimental results show that the uniaxial compressive strength of the reinforced rock mass with cross-flaws in this research is higher than that of reinforced rock mass with a single flaw. This observation can be explained by the difference in the failure modes of reinforced specimens: the reinforced rock masses with a single flaw fail due to the formation of a shear crack while reinforced rock masses with cross-flaws fail as a result of a tensile fracture or interaction between tensile fracture and shear fracture.

To provide an effective support device for dealing with the instability problems in geotechnical engineering, this paper proposes a new energy-absorbing rock bolt based on the existing energy-absorbing devices, called compression-expansion-friction bolt (abbreviated as the CEF bolt). The CEF bolt mainly consists of two parts, a variable-diameter tube and a hollow cone attached at the distal end of the bar. Controllable constant resistance and large displacement are the characteristics of the CEF bolt. The constant resistance was generated by the expansion of the tube under the compression of the cone, as well as the friction between the tube and cone. And the large displacement was provided by the sliding of the cone relative to the tube. A theoretical model was developed to predict the constant resistance by taking into account the effect of material strain-hardening. Based on the energy conservation law, the effect of shear deformation was also considered. Further, the current model was formulated as three forms to consider different conditions. The predictions of the current theoretical model were in good agreement with existing experimental data. An experimentally validated finite element model was established to study the effects of main parameters on the constant resistance of the CEF bolt, such as cone angle, tube wall thickness, friction coefficient, expansion ratio and the strain-hardening property of the tube. The results of numerical simulation and theoretical analysis show that the constant resistance was directly proportional to the tube wall thickness, friction coefficient, expansion ratio and the strain-hardening modulus of the tube. But the constant resistance increased first and then decreased with respect to the cone angle, with an inflection point existed between 5° and 10°. The shear stress distribution and the deformation of the tube were investigated as well. According to the shear stress distribution, the zone under expasion of the tube can be divided into three characteristic areas: stress concentration area, smooth transition area, and stress re-concentration area. In addition, compared with the large number of experimentally validated numerical results, the wide applicability and sufficient accuracy of the current theoretical model were verified. By taking the derivative of the current theoretical model, the location of the inflection point can be obtained. These results can provide theoretical basis for the design and field application of the CEF bolt, as well as the similar structure.

To prevent large tunnel deformations caused by the rock bursts or the squeezing ground conditions, the rock bolts used should satisfy both strength capacity and the required deformability. Currently, the energy-absorbing rock bolts, such as the cone bolt and D bolt, have been successfully utilised in the deep mining to avoid sudden tunnel collapses. However, a rigid type rock bolt (e.g. a fully grouted rock bolt) is still commonly used in civil engineering tunnels even when they are excavated under a high overburden pressure with poor geological conditions. In such cases, the rock bolt might fail in tension due to the large deformation. This paper proposed a new energy-absorbing rock bolt, which is referred to as a deformation-controlled rock bolt (DC-bolt). The performance of the proposed DC-bolt was verified by the numerical simulations using the discontinuous deformation analysis (DDA) and by the prototype laboratory tests. This study concluded that the DC-bolt possesses both the high loading capacity and the deformable capacity. Additionally, the DC-bolt can limit the rock surface movement when it reaches a certain displacement, thus, it can be a useful tunnel support for civil engineering tunnels excavated under the squeezing ground condition.

This study proposes a new analytical solution to predict displacement of a fully grouted rock bolt intersected by single rock joint. The main characteristics of the analytical model, consider the bolt profile and joint movement under pull test condition. The anchorage capacity of fully grouted bolts has been studied for many years; however, the bolt profile and its effect on bolt shear resistance are poorly understood. Investigations of load transfer between the bolt and grout indicate that the bolt profile shape and spacing play an important role in improving the shear strength between the bolt and the surrounding strata. Rock displacement is a sum of elastic part and a jump part due to the presence of joints planes. The performance of the proposed analytical model is validated by experimental method and comparison with numerical modeling. The results showed that there is a promising agreement between analytical and numerical methods. Studies indicate that the displacement rate between the bolt and the rock declines exponentially. Which is dependent on the bolt characteristics such as: rib height, rib spacing, rib width and thickness grout, material and joint properties.

Hollow grouted rock bolts are widely used to reinforce the surrounding rock mass in underground engineering, such as deep roadway of coal mines, deep tunneling. The anchorage capacity of hollow grouted rock bolts greatly influences the support quality. However, the anchorage capacity of hollow grouted rock bolt is usually destructively measured. In this paper, the grout ratio, defined as ratio of the grouted volume to the over-coring volume of anchorage, is introduced to measure the anchorage capacity. The quantitative relationship between the grout ratio and anchorage capacity of hollow grouted rock bolt by pull-out tests is established. Then, the non-destructive inspection on grout ratio of anchorage for hollow grouted rock bolts by stress wave reflection method is theoretically and experimentally investigated. By combining the two steps, a new non-destructive inspection method on anchorage quality of hollow grouted rock bolts is proposed. The results indicate that: (i) The anchorage capacity increases with the grout ratio increases up to 75% beyond which the benefits are lessened. (ii) The stress wave propagating along the bolt rod reflects and transmits at each interface impedance mismatch. The reflected signals can be used to identify the start and end of anchorage, and void part in anchorage. (iii) The results of wavelet transform analysis indicate the low-pass filters employed to reflected signals can be used to identify the start of anchorage, while the high-pass filters can identify the void part and the end of anchorage. (iv) A new supporting capacity monitoring method combined with pull-out tests and non-destructive method on hollow grouted rock bolts has been proposed, which can be used to inspect the on-site hollow grouted rock bolts in roadway on a large scale.

As rock bolts have become one of the main support systems in tunnels and underground structures, the integrity of the rock bolts affects the safety of these types of structures. The purpose of this study is the evaluation of rock bolt integrity using Fourier and wavelet transforms of guided ultrasonic waves. Rock bolt specimens with various defect ratios are embedded into a large scale concrete block in the laboratory and in the rock mass in the field. Guided waves are then generated by a piezo disk element and measured by an acoustic emission sensor. The captured signals are analyzed in the frequency domain using the Fourier transform, and in the time–frequency domain using the wavelet transform based on a Gabor wavelet. The magnitude spectrum obtained from the Fourier transform shows that a portion of high frequency content increases with an increase in the defect ratio. The peak values in the time–frequency domain represent the interval of travel time of each echo. The energy and phase velocities of the guided waves increase with the defect ratio. This study shows that the magnitude spectrum ratio and the energy and phase velocities may be indicators for the evaluation of rock bolt integrity.

This paper focuses on the grout cohesive strength and the grout friction angle of rock bolt, which have a significant influence on the pullout force and are difficult to estimate. Traditional method estimate the two parameters from the results of pull-out test conducted under different confining pressures. St. John and Van Dillen gave two approximate empirical formulas in 1984. In this text, a new method was proposed to back calculate the grout cohesive strength and the grout friction angle based on given field pullout force of rock bolt. In order to verify the method, a numerical model was built by FLAC 3D to approach the two parameters by the principle of dichotomy. The convergence result had been proved to be right by numerical pull-out tests.

High stress in the surrounding rock mass can cause serious stability problems. The applied support system used under high in situ stress condition should be able to carry high loads and also to accommodate large deformation without experiencing serious damage. This paper presents a specifically designed rock bolt for energy-absorbing applications, which can provide support for squeezing and burst-prone rocks often encountered during underground excavation in the tunneling and mining community. The bolt mainly consists of a smooth steel bar with an anchor near the bottom end of its body. The anchor is firmly fixed within a borehole using either cement grout or resin, while the smooth section of the bolt inserted in the anchor can slide in response to rock deformation once the load exceeds the pre-set capability. Static pull tests on the new bolt show that it can elongate to any expected length at a high load level, thereby absorbing a large amount of energy to maintain the stability of surrounding rock.

Friction type steel rock bolts are one of the most commonly used for reinforcing tunnels (since the late 1970s) due to their ability to sustain large rock mass displacements. In this paper, corrosion behavior of an FRS rock bolt (ASTM A607 steel) was investigated under Yucca Mountain simulated concentrated water at different concentrations (1×, 10× and 100×) and temperatures by using electrochemical measurements. Effects of individual salts, chlorides, sulfates, bicarbonates and silicates were also evaluated using the same method. Overall, the corrosion rate was an order of magnitude higher in an aerated condition than in a deaerated condition. The results from the individual salts showed the highest corrosion for the bicarbonate salts (∼300 μm/yr) and lowest for the silicate (4 μm/yr). The bicarbonate solutions accelerate the anodic dissolution of steel at higher concentrations. In the case of silicate solutions, the rock bolt shows passivity even at a higher negative potential (−925 mV) compared to other salts (∼−750 mV) with a large passive region which forms a protective layer on the specimen surface.

Energy absorbing rock bolts are used as part of rock support systems in underground constructions that are exposed to e.g., rock bursts and detonating explosives. A rock bolt capable of absorbing kinetic energy from these loads must be able to yield with the ground movements and also deform plastically over large distances, at high displacement rates. A new type of energy absorbing rock bolt has been developed and tested in laboratory. The bolt is without a casing and consists of a steel bar that has an inner ribbed-like anchorage section and an outer nut that transfers the load from the rock via a circular disc. When subjected to a dynamic load, the lengthening of the steel bar leads to a decrease in diameter whereby the adhesive bond between bar and grout is lost and the outer end of the bolt is free to yield. The rock bolt is given a very good protection from corrosion when fully grouted in cement. In a laboratory, rock bolts in concrete cylinders were subjected to free fall tests to achieve a loading velocity of 10 m/s. The tests demonstrated that the distribution of plastic strain along the length of a grouted rock bolt is not constant when dynamically loaded. The sections where plastic yielding was allowed were not fully utilized in any of the cases, opposite to that in previous static tests which show almost constant elongation of the bolts. The tests also verified that the load-carrying components of the bolt, the nut and the anchorage, are reliable when dynamically loaded. Elastic and plastic waves will start to propagate through the rock bolt as it is suddenly loaded, resulting in permanent deformation along a section of the bolt. This yield process is demonstrated through a combined graphical and numerical method.

An analytical model is proposed to predict the axial force of grouted rock bolt in the tunnelling design. The interaction mechanism of the rock bolt and the soft rock mass has been described according to their consistent displacement. Coupling and decoupling behaviors of the rock bolt around a circular tunnel have been analyzed. According to case studies, the theoretical prediction of the axial force agrees well with the in situ measured results. The installing time and the length of the rock bolt, and the deformation modulus of the rock mass are taken as study parameters to analyze the supporting behavior of the rock bolt. According to the results of the theoretical analysis, there are some conclusions as followings: (1) a lower axial force is resulted because of the delay of installing rock bolt and its supporting effect may be reduced; (2) the larger displacement is caused by the lower deformation modulus of the rock mass, and a higher axial force is resulted in the rock bolt. If the shear strength of the rock mass is not enough, the decoupling behavior will take place interior the rock mass, and the performance of rock bolt may be reduced; (3) the position of a neutral point is related with the radius of tunnel, the physical properties of the rock bolt and the rock mass. It is found that the position of the neutral point and the maximum axial force of the rock bolt installed in the soft rock may tend to be constant when its length is long enough, which means that the supporting effect of the rock bolt can not be improved significantly only by increasing the length of the rock bolt. By using this model, a way is supplied to analyze the supporting behavior of the rock bolt, and a method is provided for the quantitative evaluation of its supporting effect in NATM tunnelling.

This paper represents the result of a project conducted with developing a safe, practical and economical support system for engineering workings. In rock engineering, untensioned, fully cement-grouted rock bolts have been used for many years. However, there is only limited information about the action and the pull-out load capacity of rock bolts, and the relationship between bolt-grout or grout-rock and the influence of the grout properties on the pull-out load capacity of a rock bolt. The effect of grout properties on the ultimate bolt load capacity in a pull-out test has been investigated in order to evaluate the support effect of rock bolts. Approximately 80 laboratory rock bolt pull-out tests in basalt blocks have been carried out in order to explain and develop the relations between the grouting materials and untensioned, fully grouted rock bolts. The effects of the mechanical properties of grouting materials on the pull-out load capacity of a fully grouted bolt have been qualified and a number of empirical formulae have been developed for the calculating of the pull-out load capacity of the fully cement-grouted bolts on the basis of the shear strength, the uniaxial compressive strength of the grouting material, the bolt length, the bolt diameter, the bonding area and the curing time of the grouting material. © 2002 Elsevier Science Ltd. All rights reserved. The authors would like to thank the Research Fund of Cukurova University for supporting this research. No. FBE.2000.YL.50.

The support of rock bolt plays a significant role in maintaining the stability of a roadway in steeply inclined geological formations, yet choosing a reasonable support scheme of rock bolt in this situation is still not well understood. To reveal the force or stress distribution mechanism of anchor bolts and design an effective support scheme, an in situ anchorage experiment was conducted in a deep-buried roadway in steeply inclined rock strata. This results of this study demonstrate that in the steeply inclined rock formations, the intersection angle of a bolt length direction and the interface of rock layers have a great effect on the distribution and variation of axial force in bolts. The axial forces at the positions near the intersection of the interface and bolt increase a lot with time. In addition, the existence of no small or non-zero axial force at the end of anchor bolts indicates that much longer bolts should be used to prevent further deformation in steeply inclined rock.

In mining operations, the stability of excavations is critical to ensure the safety of production. Rock bolt support is the most effective and economical method commonly used to improve ground stability. The glass-fiber-reinforced plastic (GFRP) bolt has been developed and is becoming widely used because of its corrosion resistant character and its high tensile strength. However GFRP bolts are more susceptible to damage by blast loads and therefore it is important to understand the mechanical behaviour of GFRP bolts under dynamic loads so the support system can be better designed. In this paper we propose a new approach to describe the mechanical behaviour of fully grouted GFRP rock bolts under the combination of pre-stressed static load and blast dynamic load using the Kelvin’s solution. The method was applied to a case study and it was found that large axial and shear stresses of fully grouted GFRP rock bolts are mainly distributed in a small region near the bolt collar. It was further demonstrated that the strength of blasting, the distance between the rock bolt and the blasting point and the Young’s modulus of the rock mass and the bolt material are the most important factors having significant influences on the mechanical behaviour of GFRP bolts. For the cases examined where the bolt is within the range of 15 m from the blasting point, the influence of blast load on the axial and shear stresses of the rock bolt is significant and the dynamic load is the main factor causing the support damage. The proposed model provides a method to quantitatively evaluate the effect of blasting on the mechanical behavior of GFRP rock bolts so that a better ground support system can be designed in practical mining applications.

Rock quality is a critical design parameter of rock bolts. However, rock weathering deteriorates rock quality and adversely affects the performance of rock bolts. The objective of this study was to investigate variations in the velocity and predominant frequency of guided waves in rock bolts subjected to rock weathering. In experiments, chemical and mechanical weathering on 10 rock specimens was achieved with salty water and a slake durability test, respectively. The measured properties of weathered rock specimens were used in a numerical simulation model. Numerical studies on rock bolt surrounded by weathered rock mass were performed to simulate guided waves in the rock bolt. The results show that the group velocity and predominant frequency of the rock bolt decreased the weathering progressed. The grey relational grade value and importance score based on the random forest algorithm indicate that the group velocity and predominant frequency are significant factors sensitive to rock weathering. This study demonstrates that the group velocity and predominant frequency of guided waves can be useful tools for predicting rock weathering surrounding rock bolts.

Among the analyses of fully-grouted rock bolt in the rock mass, the effect of the end plate on the global stability of the tunnel has not been comprehensively reported. In this paper, a simplified numerical procedure for the interaction between the fully-grouted bolt and rock mass with consideration of the end plate is proposed. The rock-bolt interaction is considered at the initial and final states. At the initial state, immediately after the bolt is installed, a finite difference solution of the stresses and strains in the rock mass is proposed. At the final state, when the fictitious support pressure is negligible, another finite difference solution of the stresses along bolt as well as displacements in the rock mass is presented. To obtain the final results, a simplified iterative process is performed by repeating the finite difference solution at the final stage. The proposed procedure is verified by comparing with the commercially available finite difference software and field results. In the parametric studies, the effects of the end plate for the geological conditions with variation in rock mass quality, in-situ stress, shearing stiffness at the bolt-grout interface, and end plate stiffness are simultaneously investigated. The results indicate that, under the squeezing conditions, the failure mode of the bolt is influenced by the end plate stiffness; an optimum design for the shearing stiffness and end plate stiffness can be implemented in order to achieve an excellent supporting performance of the rock bolt.

An analytical model is developed to simulate the load-displacement performance of rock bolts under axial loading based on Modified Continuously Yielding Criterion. Such a model accounts for various surface geometries of rock bolt as well as the various confining pressures at the rock bolt to the grout interface which play significant roles in the behaviour of grouted rock bolts under axial loading. Majority of the defined parameters in the model can be measured within the laboratory environment. The proposed model is validated against the experimental results obtained from FRP and HA25 rock bolts. The good agreement between the experimental results and the analytical predictions highlights the ability of the proposed model to simulate the full load-displacement behaviour of rock bolts under axial loading.

The three-dimensional principal stress mutation of surrounding tunnel roof rocks can cause devastating disaster during excavation process. The mechanism of stress redistribution thus becomes critically important to implement the rock bolt support and enhance the safety and stability. This research focuses on the effect of rock bolt installation relative to the optimal orientation alternatives. There are deficiencies in contemporary analytical literature for that purpose. This shortcoming not only increases the construction expenditure unnecessarily but also impacts the engineering stability. The problem with 3D principal stress rotation is that uncertainty in finding the optimal bolt rock implementation during tunnelling. The stereographic projection technique that 3D is transformed to the 2D chart with the benefits of simplicity and brevity. This algorithm allows tunnelling practitioner better understanding and firmer grasp of complexity affiliated with the 3D principal stress rotation process. First we use FLAC3D to simulate the unsupported tunnelling excavation process, to produce the stress redistribution and orientation mutational data of surrounding tunnel roof rocks. We then take the modeling result to generate Stereographical Projection. The result clearly shows the progressive principal stress rotation during tunnelling, and renders the optimal rock bolt orientation.

In addressing limitations of conventional ground support measurement, a novel optical strain sensing technique is presented that is capable of measuring strain at increments as low as 0.65 mm along the entirety of an optical sensor affixed to a support element. The technique considers monitoring three sensing lengths along the profile of a fully grouted rock bolt element using a single optical fiber, which, in turn, allows the derivation of both the principle strain and principle strain direction along the bolt. A series of three experiments, which include: symmetric bending, combined axial load and bending, and double shear loading are presented and highlight the potential of the technique to capture bolt behaviour under such generalized rock bolt loading conditions. Analogized as a distributed strain rosette, the optical technique can distinguish both the coaxial and bending induced constituents of the total strain in the bolt regardless of load orientation.

When constructing an excavation, the segmental lining installed in an anisotropic rock mass is susceptible to asymmetrical pressure and substantial local instabilities. To promote liner stability in such cases, a support system that combines a segmental lining with a highly deformable expanded clay and rock bolts is proposed. A series of physical model tests were performed to study the effect of this support system. In the tests, different dip angles of the primary rock mass jointing, different thicknesses of the expanded clay and different lengths and spacings of rock bolts were considered. Variations in the internal forces and deformation of the segmental lining were recorded in each case. The distribution of the internal forces and deformation of the segmental lining were anisotropic due to the rock mass anisotropy; the maximum positive bending moment and positive deformation on the liner mainly developed in the direction normal to the stratification. By using a sufficiently thick deformable clay layer, the anisotropy of the load transferred to the segmental liner can be minimized. By adding rock bolts, the combined support system can take full advantage of the reinforcement effect of the rock bolts and the yielding effect of the deformable clay layer, and the effect of the rock bolts plus the yielding layer is greater than the sum of the two individual effects. Such a system can change the internal force and deformation distribution of the segmental lining to improve liner stability.

Rock bolts are widely employed as an effective and efficient reinforcement method in tunnel engineering. Owing to the difficulty in considering the rock breakage due to tunneling, and lack of the ability to reveal the microscopic mechanism of the interaction of bolt and rock mass properly by the conventional numerical methods, the reinforcement mechanism of bolts is not very clear. A new efficient hybrid DEM-Analytical model is proposed in this study for analysis of the deeply buried tunnels, where the DEM and analytical solutions are respectively employed to model the nearby and far fields of tunnel opening. The accuracy and applicability of the model are tested by comparing the hybrid model with analytical and full DEM model. The largest deviations of stresses and displacements for elastic problem are within 7–8% from the analytical solution, and the bond breakage regions obtained by hybrid model and full DEM model are about the same on average when the rock breakage is considered. The computational time of hybrid model is only about one-third of that of the full DEM model in this case. In addition, the variation of bolt axial forces obtained by the hybrid model is in good qualitative agreement with field measurements. The reinforcement mechanism of fully grouted passive rock bolting and the effect of supporting parameters are analyzed by the proposed hybrid model. Some useful numerical results are presented. It is found that rock bolts prevent the breakage of particle bonds immediately around them, and two transmission patterns of bond forces are observed in the nearby areas of rock bolts. Increasing the number of bolts contributes to significantly reduce the rock breakage region and bond breakage number. A threefold reduction of the number of broken bonds occurs passing from 4 to 12 rock bolts. The reinforcement of rock bolts is not obvious when the length is less than 0.5 r 0 , however, the increase of bolt length is not efficient in preventing rock breakage when the length exceeds 1.0 r 0 - 1.2 r 0 . Therefore, a reasonable efficient bolt length is suggested under the specific conditions of this study.

A rock and/or coal burst is an uncontrolled failure that releases a massive amount of kinetic energy, inducing excessive displacement of coal/rock mass. Combined support to controlling rock and/or coal dynamic failures is an essential part of the rock burst management. In the design of rock support, it is essential to consider not only the capacity of the individual elements but also their compatibility with each other and their interactions. This paper aimed to develop a novel full-scale numerical procedure to evaluate the behaviour of individual support elements and their interactions under dynamic loads using Finite Element Commercial Package ABAQUS/Explicit. A mutual interaction relationship between two indicators, namely Cable/Rock Bolt Dissipated Energy and the Steel Mesh Dissipated Energy is also developed, and an interaction diagram is provided. The proposed design concept can be used for the selection of an appropriate support system for coal burst management.

This paper modified the original structural elements of pile in FLAC3D to analyse the performance of rock bolts. For the contact surface between the bar and the binding grout, bonding behaviour was modelled using the tri-linear constitutive relation. Pull tests were used to testify the effectiveness of the modified pile elements. A satisfactory correlation was acquired between the modelling and physical results. This testified that the improved pile elements can successfully analyse the performance of bolts. Then, a numerical coal roadway was established and reinforced with the modified pile elements. After equilibrium, a non-uniform force distribution occurred along bolts. The axial force around the middle was maximal and the tensioning force at two ends of the bolt was minimal. A parametric analysis was performed to analyse the effect of relevant factors on performance of bolts. The results showed that with the bonding strength of the bolt surface raising, the neutral point shifted towards the excavation direction. Moreover, when the applied vertical stress was beyond a limit, the axial force in the bolts installed in the roadway side reduced dramatically. However, for the bolts installed in the roof, the axial force in the bolts was still raising.

Two modifications are made to enable the use of the discontinuous deformation analysis (DDA) method to investigate the sequential construction of an underground cavern complex. One modification is the “block birth and death” approach. With the aid of a boundary block searching algorithm, this approach excavates the blocks in any closed domain step by step and is verified by a DDA example. The other modification is an updating algorithm of the rock bolt called the dull-reactivating method. In this new algorithm, the coordinates of the ending endpoints of the dull rock bolts are regularly updated and the corresponding rock bolts in the sequential reinforcement region are reactivated. The initial geostress field and the sequential construction of the underground cavern complex of the Dagangshan hydropower station in Southwest China are simulated with the modified DDA method. Valuable results are obtained by analysing the displacement, stress and rock bolt force. In the construction processes, the blocks around each cavern form a bearing capacity system to prevent the block deformation, and the rock bolts jointly support the stress release as a system. At the sidewalls of the main machine building and the downstream side wall of the tail surge chamber, the blocks cut by joints with a high dip angle suffer a strong stress release, and the tension forces of the several rock bolts are so large that they should be noticed, which provides a basis for the construction guidance of the project.

This paper discusses the result of 181 pull tests of rock bolts and cable bolts in the sandstones and shales of the Sydney region. The pull tests were carried out as part of the Sydney Metro Northwest tunnelling project. The tested bolts include: end-anchored corrosion-protected steel bolts, resin-grouted fibre reinforced plastic bolts, cement-grouted multi-strand cable bolts and friction bolts. Recorded and corrected bolt stiffness characteristics are presented by bolt type, grouting agent, lithology, bolt bar material and borehole diameter. Rock-grout and grout-bar interface bond strengths are assessed for different rock mass units and bolt parameters. Common problems with typical pull test arrangements and their impact on the resulting data are discussed. Examples of issues encountered during bolt installation, such as partial grout encapsulation and insufficient mixing of resin, are presented. Recommendations are made on how some of these problems can be detected and prevented.

A comprehensive study including laboratory tests and numerical modeling was performed to investigate factors of rock bolt fracture. Experimental tests were performed to evaluate the mechanical performances of bolt ends including the thread and adjacent. The effects of installation angle, pretension and accessories (plate, washer and nut) were evaluated. It is found that a bolt installed with an oblique angle to roadway surface suffers a complex stress combination of tensioning, bending and twisting on the bolt. This complex stress is much greater than the tensile stress, leading to cracking and growth at the location of bending causing the thread to fail even though the tensile stress is less than the yield strength. Bolt thread is more vulnerable to fracture than rebar.

Referencing the design method of frame-core wall structure used in super high-rise building, a new type supporting system of grid steel frame-core tube was designed, and the structure composition and assembly method for each element were illustrated. A field test was conducted in Jinpingyan tunnel located in the area influenced by 2008 Wenchuan earthquake and the settlement of surrounding rock, internal displacement of surrounding rock, pressure between surrounding rock and primary support, internal force of support arch, axial force of rock bolts, pressure between primary support and secondary lining, and strain of secondary lining were measured to compare under two different support arches of common I-shape steel arch and new type support structure. Moreover, a 3D numerical analyses using FLAC3D software were conducted and the numerical results were compared with results of field test. It was suggested that the new type support structure played a significant role to control the large deformation of surrounding rock.

The extent of the loosening zone in tunnels excavated through columnar jointed basalts is studied with the numerical discontinuous deformation analysis (DDA) method. The structure of the rock mass and tunnel geometry are modeled on the basis of a real field case study of deep tunneling performed in such a rock mass in south west China. Slender prismatic keyblocks formed by the intersection of broadly-spaced, gently-dipping breccia layers and closely-spaced, orthogonally oriented, steeply dipping columnar joints result in a highly anisotropic rock mass structure and give rise to sliding and toppling failure modes in the sidewalls and to an excessive height of loosening zone in the roof, the geometry of which is shown to be controlled by the orientation of the steeply inclined columnar joints. Results of displacement monitoring performed during tunnel excavation in a similar rock mass with multiple point borehole extensometers confirm the numerically obtained depth of the loosening zone both in the sidewalls and the roof. We find that the height of the loosening zone in the roof as obtained with DDA is greater than would have been predicted by Terzaghi’s empirical rock load classification for “blocky” rock masses, and show that its shape and orientation are controlled by the anisotropy of the rock mass structure. Moreover, we demonstrate that dimensioning rock bolt reinforcement using well-established empirical criteria without consideration of the anisotropic nature of the rock mass may lead to un-conservative design.

The performance of two types of rock bolts, fully encapsulated rebar and D-Bolt, under combined pull and shear loading were studied in the laboratory. A new test approach was developed to apply the pull and shear loads to the rock bolt at the same time so that the bolt is displaced in a direction different from the bolt axis. Five displacing angles between 0° (pure pull) and 90° (pure shear) were employed in the tests. The test results show that the linear elastic stiffness of both the D-Bolt and the rebar bolt is mobilised quickly after a small displacement. When the displacing angle is larger than 40°, grout crushing may occur underneath the bolt shank, resulting in reduction in the stiffness of the bolt. The ultimate load of the bolts remains approximately constant no matter what the displacing angle is for both the D-Bolt and the rebar bolt. The displacement capacity of the D-Bolt, however, is dependent on the displacing angle. The ultimate displacement of a 1-m long D-Bolt section varies from 140 mm under pure pull (0°) to approximately 70 mm when the displacing angle is larger than 40°. The ultimate displacement of the rebar slightly increases from 29 mm under pure pull to 53 mm under pure shear. In general, the displacement capacity of the D-Bolt is larger than that of the rebar bolt. It is approximately 3.5 times the rebar under pure pull and 50% higher than rebar under pure shear. The test results show that the displacing angle of the bolt is larger than its loading angle, which is also confirmed by the analytical solutions.

Rock joints are prone to open and fail when subject to severe dynamic extension, which might result in spalling of surrounding rock and collapse of underground openings. However, the understanding of dynamic response of rock joints and underground openings subject to tensile stress wave is still at its infancy. To investigate the effect of tensile wave on the response, stability and supporting of underground openings in jointed rock mass, numerical modelling was carried out herein with the DEM-based universal distinct element code (UDEC) after its validation using explosion testing measurements. Results showed that joint tensile strength has significant effects on stress wave propagation if it is lower than the amplitude of tensile stress wave in the shallow rock mass, where joint opening will occur, no significant portion of wave energy could transmit through the joint, and the transmission coefficient for tension-first wave is lower than that for compression-first wave. The buried depth of underground opening and joint properties including stiffness, spacing and dip angle as well as crossing angle between joint sets could significantly influence stress wave propagation and dynamic responses of underground openings under dynamic extension. In addition, installation of rock bolts with appropriate number and length and shotcreting with sufficient thickness could effectively reinforce the surrounding rock and reduce the area of disturbed zones when against dynamic extension. The findings in this study could be of great significance for the design, supporting and stability evaluation of underground openings.

Rock bolts are a crucial part of underground infrastructure support; however, current methods to locate and record their positions are manual, time consuming and generally incomplete. This paper describes an effective method to automatically locate supporting rock bolts from a 3D laser scanned point cloud. The proposed method utilises a machine learning classifier combined with point descriptors based on neighbourhood properties to classify all data points as either ‘bolt’ or ‘not-bolt’ before using the Density Based Spatial Clustering of Applications with Noise (DBSCAN) algorithm to divide the results into candidate bolt objects. The centroids of these objects are then computed and output as simple georeferenced 3D coordinates to be used by surveyors, mine managers and automated machines. Two classifiers were tested, a random forest and a shallow neural network, with the neural network providing the more accurate results. Alongside the different classifiers, different input feature types were also examined, including the eigenvalue based geometric features popular in the remote sensing community and the point histogram based features more common in the mobile robotics community. It was found that a combination of both feature sets provided the strongest results. The obtained precision and recall scores were 0.59 and 0.70 for the individual laser points and 0.93 and 0.86 for the bolt objects. This demonstrates that the model is robust to noise and misclassifications, as the bolt is still detected even if edge points are misclassified, provided that there are enough correct points to form a cluster. In some cases, the model can detect bolts which are not visible to the human interpreter.

Passive bolting is used to stabilise unstable rock blocks in surface and underground structures due to the various advantages it offers. Despite its use, the design phase still presents aspects of considerable complexity because the fact that the load of the bolt and therefore, its static action depends on its interaction with the block and the stable rock. In the present work, a mathematical model was developed which is capable of directly calculating the stabilisation forces as a function of the characteristic parameters of the bolt and of its interaction with the rock. This discussion is based on a simplified hypothesis of bolt behaviour, which provides negligible errors, and on the observation that the critical point is positioned at the intersection of the bolt with one of the lateral surfaces that separate it from the portion of stable rock. The formulation of the stabilisation forces obtained made it possible to evaluate the static contribution of each single bolt to the stability of the rock block, by varying the diameter of the steel bar and then designing the bolting operation to achieve acceptable stability conditions for the rock block. The application of stabilising equations to a real case, for which the results of load tests on bolt tests were available, allowed us to outline steps to be taken in the bolt design process.

An axial splitting energy absorber for rock bolts was designed to adapt to the large deformation of surrounding rocks in deep roadways. This paper presents an experimental investigation of the mechanical behavior of the splitting energy absorbers with different parameters. The effects of materials, diameter-to-thickness ratio, prefabricated cuts, induced splitting dies, anti-curling splitting conditions, and oblique loading on the performance of the components were studied. The test results indicated that the splitting process of the component made of Q235 steel with four prefabricated cuts, is more stable; the smaller the diameter-to-thickness ratio, the higher the constant axial load level of the component; the anti-curling condition can only increase the load level of the components by 6%; the components have the advantages of “strong deformation sensitivity and weak load sensitivity” when subjected to oblique loading. Eventually, the parameters of the energy absorber were optimized for the rock bolts, i.e., the components have an inner diameter of 26 mm, a wall thickness of 4 mm, and four prefabricated cuts with a free-curling splitting mode. The novel anchor bolts exhibited stable energy absorption behaviors when subjected to static and dynamic loading. Moreover, the energy absorbers increased the stability, thereby yielding a rebar stroke of 40–200 mm. More importantly, serial research work on energy-absorbing components with various parameters will establish the experimental foundation for the application of components in various mining environments.

Most retreating roadways in extra-thick coal seams, are affected by multiple mining activities under fully-mechanized top caving mining conditions. Surrounding rock deformation and stress evolution of the roadways are complex under intensive mining activities due to interlayers occurrence in top thick coal roof developed in fissures and fractures in coal seams, and large cross-section of the roadways. If no appropriate support techniques and optimization measures are taken under such mining conditions, consequences such as severe bedding separation, large deformation of surrounding rock, and large-area top caving disaster, can occur. Therefore, the investigation of surrounding rock deformation and failure mechanism of retreating roadways and the control technique for extra-thick coal seams under fully-mechanized top caving and intensive mining conditions, is the key for stability control for this type of roadway. Various research methods were adopted comprehensively in this study including field investigation, monitoring and statistics, experimental tests, physical similarity simulation, and field testing. Taking the return airway # 5110 in the main mining coal seam # 3–5 with the thickness of 14.7 m of Tashan coal mine, Datong Coal Mine Group as the case, failure characteristics of roof supporting structures of the return airway # 5110 undergoing retreating mining of the upper coal panel # 8109 and the current mining panel # 8110 were investigated. By employing bedding separation acquisition monitoring device and force-meters, surrounding rock deformation, carrying the load of rock bolts and cables, and bedding separation, were monitored and analyzed. The monitoring was done by the surrounding deformed rock where loading characteristics of the supporting structures in the roadway # 5110 were obtained under intensive mining conditions for the extra-thick coal seam. Stress evolution during the process from roadway stability to instability was characterized experimentally under different lateral pressure conditions by using physical similarity simulation. On this basis, the single mechanical yieldable prop with high constant resistance was proposed to support the surrounding rock of retreating roadways in extra-thick coal seams. The analysis was carried on the structure and on the working mechanisms of the prop set where its working resistance under the height of 3 m, support patterns, and support techniques, and then field tests were conducted in the retreating roadway # 5110. The typical monitoring results showed that the contraction area of the trial roadway ahead of the working face decreases significantly, and that the largest displacement between two coal sides is no more than 428 mm (the largest roof subsidence is no more than 663 mm). The test results show that good control effect on surrounding rock deformation and roof support has been achieved. The study can be applied for the entire Tashan coal mine, and has significant implication for engineering projects with similar geological and mining conditions.

The time dependent viscoelastic model is extended so that excavation lining process of tunnel can be taken into account, with special attention on the analysis of the excavation damage zone around a circular tunnel under hydrostatic condition. Combining the sliding crack model and equivalent crack method, an analytical model for the evolution of fractured zone in viscoelastic surrounding rock mass is proposed. The whole construction process of tunnel is divided into three stages, in which the stress and displacement are considered separately. Then, the time dependent evolution of EDZ is implemented in a short computation algorithm, in which both extent and direction of fractured zones are determined. It is key important for the settlement of rock bolt. Through sensitivity analysis of parameters, the dependence of EDZ on time, excavation velocity, structure of heterogeneous rock mass, lining supporting system, initial geostress and the radius of tunnel are determined. Because of the time dependence of rock mass, the failure and deformation behavior continue increasing, even if the excavation process finished. In addition, the displacement after lining is set up is far more than the one at the moment of excavation just finished.

Rock bolts used for the reinforcement of underground mines, tunnels and nuclear waste repositories are made up of low and medium carbon steels, and high strength low alloy steels. Typical rock bolt systems used for rock reinforcement are mechanically anchored bolts, grout anchored, frictional rock stabilizers and strand anchors. For nuclear waste repository sites such as Yucca Mountain (YM) as well as for mines and tunnels, in addition to mechanical properties, corrosion properties are also important due to potential seepage of water through the fractures or pores in the rock. During temporary rock support period of 50–100 years, the temperature of the tunnel at YM should be maintained at ambient conditions. For any reason if the rock bolts are exposed to YM waters and high temperatures in the tunnel then there is a chance of corrosion of steel rock bolts. In this study an attempt was made to study the corrosion properties of various potential rock bolts for YM tunnel support via the aid of electrochemical corrosion testing. At ambient temperature (25 °C) all the rock bolts that were studied showed good corrosion resistance in these waters. At higher temperatures, 60 °C and 90 °C, corrosion resistance of rock bolts decreased, but due to special stress relief heat treatment of one of the frictional rock stabilizers (Swellex Mn 24) the corrosion rates were lower than all other tested rock bolts. Note: Swellex, Split set and Williams are the proprietary names of Atlas Copco, International Roll Forms, Inc. and Williams Form Engineering Corp, respectively.

The Siah Bishe pumped storage project is now under construction. The investigation results, the stability analysis and the support concept for the displacement chamber are discussed in this paper. The geometry of the transversal section and optimum maintenance system have been designed with respect to the unique geological and shear zone conditions encountered in Siah Bishe. The displacement chamber is arranged in sedimentary and volcanic rocks of Triasic and Jurasic ages. Length of the chamber is 50 m and the height is 10 m with relatively north–south trend. The results of geo-engineering surveys and statistical joint-graphics of stereograms reveal at least three discontinuity systems in the rock mass. The transversal section along the chamber was studied to analyze the stability of rock wedges inside the chamber. The frequency and width of bedding parallel shear zones determined the final support concept. Special attention was paid to the design of the transversal section so as to minimize the stress concentration in the critical regions of the wall, bottom of the chamber and corners. The results of analyses show that the chamber is stable in static and dynamic condition with 20 cm thickness of reinforced shotcrete, fully bonded rock bolt with 35 mm diameter, 3.5 m length and steel frame with 82 kg/m I section with 0.85 m spacing.

The stability analysis of large-span tailrace bifurcation tunnels with complex 3D geometries and the associated challenges are presented in this paper. A new nonlinear reinforcement analysis method is firstly proposed for the design of reinforcement systems, i.e. steel-reinforced concrete lining and rock bolts, for large-span tailrace tunnel bifurcations with complex 3D geometries. The proposed method is then implemented into FLAC3D using the FISH language to calculate the reinforcement parameters and the resultant nonlinear stress distributions. After that, it is validated by comparing the concrete lining stresses and reinforcement obtained by the proposed method with the analytical solutions for the ideal case of a circular tunnel in an infinite elastic body. Finally, the new method is applied to calculate the distribution of deformations, stresses, plastic zones and design the reinforcement system for the large-span tailrace tunnel bifurcation in the Guandi Hydropower Station during the construction, operating and maintenance periods, respectively. The calculated reinforcement parameters successfully guided the construction of the tunnel. The proposed nonlinear reinforcement analysis method may also be beneficial to the design and construction of other large-span tunnel bifurcations with complicated irregular cross sections in similar hydropower projects currently ongoing in southwest China.

The estimation of support requirements to stabilize underground structures is of prime importance for rational design of these structures. The characterizing parameters of rock mass may vary with depth. Determination of these parameters by drilled cores and Ground-Probing-Radar (GPR) is difficult and expensive due to anisotropy of rock mass. Laboratory testing is also expensive. Also the in situ conditions are difficult to simulate in the laboratory. The designer is thus resorting to empirical methods and analytical methods to determine these parameters. Often, the analytical techniques may mesmerize the designer to feel the problem and its solution on the screen of the computer. In this paper, an attempt has been made to develop algorithm based on Block Theory with geological information & mechanical properties of rock for determining the rock pressure. Limitations of this technique are number of joint sets not less than three and width of the opening up to 25 m. The algorithm determines all the wedges formed at a time by 3, 4, 5, 6, …, n joint planes with excavation plane responsible for manifestation of rock pressure at roof/wall. All the permutations and combinations for wedge formation can be considered in this respect. Rock pressure for design is determined for reinforcement of the underground openings. Spacing of rock bolts is found out as an additional feature. The alignment of the opening for optimal reinforcement can also be determined. Case history of Tehri Power House, India is taken up for analysis. The empirical correlations developed by Goel (1994) are used for comparative study. It was found that no appreciable rock pressure was developed at walls. Roof pressure is determined to be 140 kPa, which is almost same as observed. It is thus established that block theory may be applicable for design criterion up to depth of 500 m.

Today, there is a great need for larger underground spaces for various purposes and hence, construction of new metro tunnels has become a necessity to meet the demand in urban life in spite of certain ground related difficulties such as fault zones, altered and fractured rock mass and ground water. This study has aimed at investigating the risky areas around a shallow metro tunnel in weak, faulted rocks and determining the effects of tunnel behavior on the structures on ground surface. Therefore, an attempt has been made to determine the risky areas on the line of the tunnel by field observations, laboratory work and computer modeling. Later, the data obtained from computer models have been compared to which obtained from in situ measurements. The results from modeling and in situ measurements were interpreted considering the current status of superstructure and the differences between pre- and post-excavation states in the ground. Finally the data obtained from the modeling analysis and measurements provided the necessity of strengthening the already used support system for the safety of the buildings on surface. Shortening the application ranges of the rock bolts, use of face nails with application of umbrella arc and jetgrout methods are among the precautions to be taken. (C) 2012 Elsevier Ltd. All rights reserved.

The assessment of the interaction between a new tunnel and existing structures is an important issue in urban areas. In this study, the effect of tunnelling on the existing support system (i.e. shotcrete lining and rock bolts) of an adjacent tunnel is firstly investigated using ABAQUS and TUNNEL3D through full three-dimensional (3D) finite element calculations coupled with elasto-plastic material models, which takes into account the tunnelling procedure, the interaction between the shotcrete lining and rock mass, the interaction between the rock bolts and rock mass, and the elasto-plastic behaviour of the rock mass, the shotcrete lining and the rock bolts. Then, on the basis of the calculated results, it is concluded that the driving of the new tunnel significantly affects the existing support system when the advancing tunnel face passes the existing support system and is minor when the face is far from it. Moreover, the support system in the side of the existing tunnel closest to the new tunnel is more significantly affected than that on the side opposite to the new tunnel. It is also found that in a region such as Sydney with relatively high horizontal regional stresses, the driving of the new tunnel will not cause considerable adverse effects on the existing support system, if the new tunnel is driven horizontally parallel to the existing tunnel with a sufficient separation, since both the tensile stress in the existing shotcrete lining in the lateral sides of the preceding tunnel and the compressive stress at the crown decrease although noticeable tensile stress increments are observed on some parts of the existing rock bolts. Finally, it is pointed out that the effects of tunnelling on the existing support system strongly depend on the position between the original and new tunnels. In terms of the stress increments on the existing support system, especially the maximum tensile stress increments on the existing shotcrete lining, the driving of the new tunnel causes increasingly adverse effects on the existing support system in a sequence of: (i) horizontally parallel tunnels with a separation of 30 m; (ii) horizontally parallel tunnels with a separation of 20 m; (iii) staggered tunnels with a separation of 30 m; (iv) vertically alignment tunnels; and (v) staggered tunnels with a separation of 20 m in the cases investigated in this study. For the relatively high regional stresses in the Sydney region, the obtained results qualitatively agree with other’s published observations from the construction of closely parallel subway tunnels.

The Rock Mass Rating method (RMR-method) is suitable only for initial estimates of the rock mass strength. The Tunnelling Quality Method (Q-method) provides rock bolt spacings and shotcrete thickness, but neglects the effects of bolt lengths and tunnel size on the support pressure. Numerical analysis of the stress situation around a tunnel shows that support pressures are up to fourteen times those estimated by the Q-method. The analysis allows optimization of bolt lengths and spacings required to control the reduction in tunnel area. The latter is an objective-oriented criterion that makes more sense than a stability criterion. The numerical results are for an idealized situation—i.e., uniform, elastic-perfectly plastic rock (Mohr Coulomb with no strain softening), with no explicit joints; idealized rock bolts; and no time dependence. Comparison with field results should be made before generally applicable conclusions can be drawn.

A number of different Water and Frost Protection Systems, or so-called inner linings, have during the years been introduced and tested in Norwegian road tunnels. At present a limited number of such systems have been approved by the Norwegian Public Roads Administration (NPRA). These systems will all be described in brief. A system based on pre-cast concrete segments will be described in details, including the installation procedure. Experience and limitations associated with this system will also be discussed. A basic design concept for using an inner lining as water and frost protection is that there is no interaction between the inner lining and the rock support. The rock support practise in Norwegian traffic tunnels will also be briefly described in this paper. Sprayed concrete and rock bolts are the main support measures. In-situ cast-in-place concrete lining is only used locally to solve severe stability problems.

Some of the supports in the permanent access tunnel of the Jiulongkou Coal Mine had failed and large deformations were observed. This had a serious impact on the safety and production of the mine. Damaged tunnels were usually repaired using steel sets or timbers in the past. Based on analysis, a new support measure was adopted using bolt, mesh and shotcrete. Repair work carried out in a 70-m-long tunnel showed that the supporting approach and techniques were satisfactory and successful.

This paper presents a theoretical method predicting the development and progress of plastic zones around tunnels in soft rocks and determining the loosening pressure which has to be taken by rock bolts or linings. Influence effects of the mechanical properties of soft rocks on the loosening pressure are examined and discussed. The judgment method used to assess the extent of rock failure is presented according to the ground characteristic curve. The ground characteristic curve shows a lowest point that refers to the minimum support pressure and the controlled tunnel wall strain when the competency factor (i.e. ratio of compressive strength to the initial in-situ pressure) is smaller than a critical value, which can be estimated from the proposed theoretical equations.

The results of laboratory experiments on changes in the mechanical properties of various types of rock at low temperature (to -160°C), and the results of thermal shock tests, and freezing and thawing tests (from 15°C to -45°C, up to 300 temperature-change cycles) are described. Thermal conductivities and the expansion (shrinkage) rate of various rocks at temperatures from +15°C to -160°C are measured and the temperature-dependent thermal characteristics are examined. Engineering problems for cryogenic storage —e.g., the critical freezing temperature of seepage water flow into the caverns, durability of rock bolt and grout materials under low temperature—are examined in some fundamental experiments.

Field tests and laboratory investigations have been carried out to evaluate the performance of grouted rock bolts subjected to close proximity blasting. The investigations comprised pull-out testing of bolts and vibration measurements on both the rock and the bolts. Laboratory work comprised compressive and flexural testing of the grout and analysis of polished thin sections of specimens consisting bolt, grout and rock. There is no observable decrease in pull-out strength for bolts installed close to the blast (3.4 m), compared to more distant installed bolts (22.0 m). Furthermore, there is no difference in pull-out strength for the bolts with the freshest grout compared to bolts with cured grout, when subjected to blast loadings. Anchorage length of the bolts was 125 mm, which gave an average pull-out strength of 140 kN, or 93 % of steel yielding load. Analysis of polished thin sections of the rock, grout and bolt does not show any variation in crack pattern or crack frequency between bolts with varying distance to the blast or varying curing of the grout when subjected to the blast loadings. Re-pulling of previously pulled bolts shows a healing effect of the grout. None of the tests carried out indicates any decrease of the bolt /grout performance. It can be concluded that fully grouted bolts can be used on or close to the face.

The Gjovik Olympic Mountain Hall has a rock cover of 25–50 m and a span of 61 m. This is by far the largest span ever used for a cavern that is open for public use. The paper reviews the development of tunnelling and the use of the underground in Norway as a background for the decision to build this recordbreaking cavern. Experience from the excavation of very large mining chambers was of crucial importance. The Norwegian approach to the support of tunnels and underground openings is discussed. Excavation and support of large caverns such as hydropower facilities, oil storages and sports halls are described. The support system for the Gjovik Olympic Mountain Hall is described in detail, and measurements performed on rock bolts and in the shotcrete are discussed. There are clear indications that the total stability of the cavern is primarily a result of favourable horizontal stresses. The majority of the 6-m-long and 12-m-long grouted rock bolts seem to carry very low loads over only short lengths of the bolts.

A semi-empirical method has been developed for the design of support systems consisting of rock bolts/anchors, shortcrete/fibre-reinforced shortcrete, steel ribs and grouted arches to support tunnels and caverns, especially in poor rock conditions. The empirical relations for mobilizing factors of shortcrete and bolts/anchors have been deduced on the conservative side from extensive data of the Norwegian Geotechnical Institute (NGI) tables and charts for support systems presented by Barton et al. (1974). There is a reasonably acceptable correlation between the empirical approach and the proposed theory of estimating the support capacity, except for abnormal cases. The proposed method is simple to apply in the field. Special designs have been suggested for complex geological conditions for which the assumptions are not valid, such as in thick shear zones, etc.