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The effects of reinforcement system on pillars were tested in laboratory, using three types of pillars with different strengths. The tests were performed using the UCS machine, to test pillar without reinforcement, pillar with rock bolt reinforcement, pillar with shotcrete reinforcement and pillar with the combination of both rock bolt and shotcrete reinforcement. Uniaxial compressive strength (UCS) testing aims to determine the effects of the reinforcement system on pillar strength. The results of this study indicate that the reinforcement system on high strength pillars causes a strength increase of 14.93% on pillar with rock bolt reinforcement, 21.45% on pillar with shotcrete reinforcement and 34.67% on pillar with combination of rock bolt and shotcrete reinforcement. On medium strength pillars, reinforcement installation shows a strength increase of 16.27% on pillar with reinforced rock bolt, 19.83% on pillar with reinforced shotcrete and 44.40% on pillar with combination of rock bolt and shotcrete reinforcement. Likewise, on low strength pillars, reinforcement installation causes a strength increase of 13.13% on pillar with reinforced rock bolt, 36.21% on pillar with reinforced shotcrete and 53.85% on pillar with combination of rock bolt and shotcrete reinforcement. The results of laboratory testing and numerical modeling indicate that the increase in strength occurs because the horizontal displacement on the surface of the pillar wall is detained by shotcrete and faceplate on rock bolt, so that the pillar seems to have confining pressure throughout the pillar wall surface, which is called as equivalent confining pressure.

The research has proved that rock mass stability can be ensured using steel resin rock bolting including rods, fast-setting resin capsules, bearing plates, strapping and tension nuts. Interaction of these structural components with roof rocks and sidewalls involves obligatory displacement of rocks. The article presents the studies into applicability of steel resin rock bolting in difficult geological conditions in coal mines in Kuzbass. The steel resin rock bolting is considered in this case as the method of ‘active’ ground control. The ‘active’ ground control consists in the capacity of steel resin rock bolts to alter main strength characteristics of rocks, to increase their resistance and stability, and, thereby, to ensure efficiency of roof support. It is found that horizontal and vertical stresses in underground excavations depend on the mining depth. It is proved that rock bolts installed in the roof rocks can prevent or confine rock fracture.

Most of the surrounding rocks of offshore liquefied petroleum gas (LPG) reservoirs occur in jointed rock mass. This paper is aimed at the reinforcement of the layered composite rock masses. First, the structural mechanical model of the elastic deformation and completely plastic deformation for the bolt was constructed. Then, the equations of axial force, shear force, and bending moment in the cross section of the bolt were obtained by the force‐method canonical equation. The shear‐slip resistance effect of the bolt was decomposed into the friction effect and the dowel effect. In addition, the quantitative relationship between shear resistance and the characteristic angles (anchoring angle, friction angle, dilatancy angle, and deflection angle) and the geometrical and physical parameters of the bolt were derived. Finally, the influence of each characteristic angle on the mechanical properties of the bolt was analyzed in details and the model was verified. The analysis results showed that the laterally constrained load of the rock stratum on the bolt can be characterized by a power function, but the optimal power should be no less than 10. The dominant position of either the friction effect or the dowel effect was changed with the matching of different characteristic angles. The increase of the friction angle and dilatancy angle of the joint surface of the rock is beneficial to improve the lateral shear resistance of the bolt. The effect of the deflection angle on the lateral shear resistance of the bolt was affected by the anchoring angle. The optimal anchoring angle of the bolt should be in the range of 50°‐60°. The results in this study have important theoretical significance to guide the offshore anchoring engineering of layered composite rock masses.

At present, the suspended monorail systems constitute a very common means of transportation in the Polish hard coal mines. The main advantages of the suspended monorail include the independence of the route from the working floor surface irregularities and the possibility to transport cargo of significant mass and size. The masses and dimensions of machines and devices transported via monorail have increased considerably in recent times. This particularly concerns the transport of longwall system elements. In Poland, the maximum speed of suspended monorail travel is 2 m/s. Due to the fact that preparations are currently underway to increase the maximum speed above 2 m/s, it is necessary to inspect what influence it will have on work safety and mining support stability. Current operational experience and tests have shown that dynamic loads induced by the suspended monorail transportation have a significant influence on the roadway support stability, working protection durability and on the monorail operators. This is particularly true during the emergency braking of a suspended monorail by means of a braking trolley, where the overloads reach 3g. Bench tests of the selected steel arch and rock bolt support elements utilised in the Polish hard coal mines were conducted in order to determine the resistance of steel arch and rock bolt supports to static and dynamic loads. The article presents the results of the tests conducted on a steel arch support in the form of the sliding joints of an ŁP/V29 yielding roadway support, which is commonly employed in the Polish hard coal mines. Tests of elements of the threaded bolts with trapezoidal threads over the entire rod length were conducted as well. The conducted strength tests of steel arch and rock bolt support elements under static and dynamic loading have shown that dynamic loading has decisive influence on the support’s retaining of its stability. Support element stability decreases along with the increase of the impact velocity. This concerns both the steel arch support and the rock bolt support.

Grouted rock bolts represent one of the most used elements for rock mass stabilization and reinforcement and the grouting quality has a crucial role in the load transfer mechanism. At the same time, the grouting quality as well as the grouting procedures are the least controlled in practice. This paper deals with the non-destructive investigation of grouting percentage through an analysis of the rock bolt&rsquo, s natural frequencies after applying an artificial longitudinal impulse to its head by using a soft-steel hammer as a generator. A series of laboratory models, with different positions and percentages of the grouted section, simulating grouting defects, were tested. A comprehensive statistical analysis was conducted and a high correlation between the grouting percentage and the first three natural frequencies of rock bolt models has been established. After validation of FEM numerical models based on experimentally obtained values, a further analysis includes consideration of grout stiffness variation and its impact on natural frequencies of rock bolt.

Usually, in a fully grouted rock bolt pullout test the load-displacement curve of the rock bolt head is recorded. This paper presents an analytical method to use this curve for determining the bond (bolt-grout and grout-rock interface) shear strength parameters. For this purpose, the fully grouted rock bolt interaction with grout and surrounding rock in the pullout test is investigated and the load-displacement curve of the bolt head (beginning of the bonded section) is obtained analytically. For modeling the bolt-grout interface behavior a distribution of the shear stress along the fully grouted rock bolt by consideration of bolt shank failure is used. In this regard, different stages including complete bonding, partial decoupling, decoupling with the residual shear strength and complete decoupling are considered. With increasing the applied load, two possible cases involving the rock bolt complete pullout and bolt shank yielding are taken into account. Based on the presented analytical method, the obtained bolt head load-displacement curve can be compared with the one recorded in the pullout test. With this, the relevance of selected shear strength parameters compared to real parameters can be assessed. A flowchart for determining the bolt bond shear strength parameters is presented using the trial and error method (coded in Matlab). The proposed solution is used to determine two experimental pullout shear strength parameters. The results show good agreement between predicted and calculated load-displacement curves.

This study aims to evaluate the sensing performance of glass fiber-reinforced polymer (GFRP) rock bolt sensors instrumented with strain gauges for monitoring rockslides. Experimental studies are conducted with four different types of GFRP rock bolt sensors and concrete blocks having central holes and two shear joints. Two GFRP rock bolt sensors are inserted into holes and then fixed in concrete blocks with cement grout and soil, respectively. The other two are coated with heat-shrink tubes to protect strain gauges and wires, which are then fixed in concrete blocks with cement grout and soil, respectively. Double shear tests are performed to produce shear deformations of GFRP rock bolt sensors, and then strain change with shear displacement is monitored. The results manifest that the variation in strain with shear displacement is more sensitive in the GFRP rock bolt sensor fixed with soil than with cement grout. Also, strain gauge wires in the GFRP rock bolt sensor fixed with cement grout are broken earlier than with soil. Furthermore, it is confirmed that the heat-shrink tube effectively protects strain gauges and wires, so that GFRP rock bolt sensors coated with heat-shrink tubes work for a longer time than the uncoated sensors. The present study shows that the GFRP rock bolt sensor can be useful for monitoring rock slope failure.

Rock joints will undergo a sequence of cyclic shearing loadings during a seismic event. However, the effect of cyclic shear loading on the energy-absorbing rock bolts has never been studied before. Laboratory shear experiments were carried out to study the shear behaviour of rock joints reinforced by the energy-absorbing rock bolts under cyclic loading condition. The results illustrated that the support effect of the energy-absorbing rock bolts was very small after the first cycles in the cyclic shear experiments. In the case of small cyclic distances, the shear resistance of the energy-absorbing rock bolts will gradually recover after the shear displacement has exceeded the cyclic distance in the subsequent shear experiment after 5 cycles. In the case of large cyclic distances, no recovery of shear resistance was found in the subsequent shear experiment, indicating that the energy-absorbing rock bolts had completely lost its supporting role after cyclic shear loading. A new index of shear energy loss ratio (SELR) was proposed to evaluate the shear behaviour of energy-absorbing rock bolt and rock joint under cyclic shear loading condition. The results showed that the SELR of rock joints was commonly less than 20%. However, the SELR of rock bolts could reach nearly 100% when the cyclic distance was larger than 8 mm. When the cyclic distance was 4 mm or 6 mm, the SELR of the fully encapsulated rock bolts almost reached 100%. However, the SELR of the energy-absorbing rock bolts were located in the range of 50–80% for the same condition. The results indicated that the shear behaviour of a rock bolt inserted in a rock joint was strongly influenced by cyclic shear loading. The shear performance of the energy-absorbing rock bolts was better than the fully encapsulated rock bolts under cyclic shear loading conditions.

Anchoring mechanism and failure characteristics of composite soft rock with weak interface usually exhibit remarkable difference from those in single rock mass. In order to fully understand the reinforcement mechanism of composite soft roof in western mining area of China, a mechanical model of composite soft rock with weak interface and rock bolt which considering the transverse shear sliding between different rock layers was established firstly. The anchoring effect was quantified by a factor defined as anchoring effect coefficient and its evolution equation was further deduced based on the deformation relationship and homogenized distribution assumption of stress acting on composite structure. Meanwhile, the numerical simulation model of composite soft rock with shear joint was prompted by finite element method. Then detailed analysis were carried out for the deformation features, stress distribution and failure behavior of rock mass and rock bolt near the joint under transverse load. The theoretical result indicates that the anchoring effect of rock-bolt through weak joint changes with the working status of rock mass and closely relates with the physical and geometric parameters of rock mass and rock bolt. From the numerical results, the bending deformation of rock bolt accurately characterized by Doseresp model is mainly concentrated between two plastic hinges near the shear joint. The maximum tensile and compression stresses distribute in the plastic hinge. However, the maximum shear stress appears at the positions of joint surface. The failure zones of composite rock are produced firstly at the joint surface due to the reaction of rock bolt. The above results laid a theoretical and computational foundation for further study of anchorage failure in composite soft rock.

The present study focuses on the load transmission law along the rock bolt and the efficacy for different anchored conditions. Shear displacement along the joint face is considered on the basis of traditional rock bolt stress calculation method with regard to joint open displacement. Combining the elastic foundation beam model and the rock bolt pull-out model, an analytical solution for the stress distribution along the rock bolt through a single joint face is derived. The performance of the proposed analytical solution is compared with an experimental result and a previous study. In addition, joint displacement, joint location, and dip angle of the joint face are set as variables to further explore the corresponding anchored effect. The results indicate that larger joint displacement along the joint face tend to cause a distinct increase of the shear and compressive stress along the rock bolt. When the joint face moves inward, the interaction force between the rock bolt and the surrounding rock mass decreases, suggesting that the strength of the rock bolt is not taken full use. Moreover, for a certain joint displacement and joint location, larger anchored angle between the rock bolt and the joint face contributes to better reinforcement effect.

Recently, conventional tunnelling methods (or the drill & blasting tunneling methods, the new Austrian tunneling methods) have been frequently applied to tunnels/caverns under severe geological conditions. Under such adverse conditions, it is a concern that tunnel support materials might fail because of the huge ground pressure and large tunnel deformation. As a result, the tunnel may lose its stability due to decreased effectiveness of tunnel supports. Therefore, essential components of tunnel supports (e.g., rock bolts) are becoming more important. According to the previous studies, the interface behaviour between the rock bolt and the bond material (i.e., the debonding process along the interface and crack propagation in the bond material) is recognised as one of the most important factors significantly affecting the reinforcement effects of rock bolts. However, there is little information available on the detailed interface behaviour. This research aims to investigate the interface behaviour via both laboratory tests (shear test) and numerical simulations using the discontinuous deformation analysis (DDA). As a results, it can be found that the DDA is an effective tool reproducing the accurate interface behaviour, and therefore the simulation results enable us to better understand the bolt-grout interface behaviour in detail. Based on the accurate DDA rock bolt model, a new energy-absorbing rock bolt was designed. This rock bolt has high loading capacity and large deformability. Moreover, it can control the ultimate tunnel deformation utilizing the specially designed anchor. By carrying out the DDA-based simulations of pull-out tests and tunnel excavation, the reinforcement effect of the newly-developed rock bolt was investigated. As a result, it can be concluded that the new rock bolt has a significant advantage over the widely-used fully grouted rock bolt when the large tunnel displacement occurs. Doctor of Philosophy

Dynamic and static coupling loads (DSLs) are one of the most common stress environments in underground engineering. As the depth of a roadway increases over the life of a mine, the static load of the ground stress field increase multiplies, and the cyclic operation at the working face releases a large amount of dynamic energy. Therefore, deep roadways easily induce dynamic disasters during production. In this paper, a deep roadway numerical model was built with FLAC3D to test the deep roadway under DSLs and was simulated with 16 different support designs. The ground stability in each support condition was examined and compared in terms of the ground deformation and scope of failure. The underlying support mechanism was further analyzed with numerical modeling in view of the deformation in the surrounding rock mass induced by variations in the support parameters. The results show that shortening the bolt spacing is an effective measure to control the deformation of surrounding rock whatever DSLs or static load. Under static load, the larger the anchoring length is, the more stable the surrounding rock is. Under DSLs, end grouting length (S = 600 mm) and full grouting length (S = 1800 mm) can effectively control the deformation of surrounding rocks and enhance the stability of surrounding rocks. The results contribute to the design of supports in the field of underground coal mines and provide a basis for determining the reasonable support scheme for roadways.

In the underground mines of the Legnica–Głogów Copper District (LGOM) the main way to protect the room excavation is the use of a rock bolt support. For many years, it has proven to be an efficient security measure in excavations which met all safety standards and requirements. The article presents the consumption of the rock bolt support in the Mining Department “Polkowice–Sieroszowice” in the years 2010–2015 as well as the number of bolt supports that were used to secure the excavations. In addition, it shows the percentage of bolt supports that were used to conduct rebuilding work and cover the surface of exposed roofs. One of the factors contributing to the loss of the functionality of bolt supports is corrosion whose occurrence may lead directly to a reduction in the diameter of rock bolt support parts, in particular rods, bearing plates and nuts. The phenomenon of the corrosion of the bolt support and its elements in underground mining is an extremely common phenomenon due to the favorable conditions for its development in mines, namely high temperature and humidity, as well as the presence of highly aggressive water. This involves primarily a decrease in the capacity of bolt support construction, which entails the need for its strengthening, and often the need to perform the reconstruction of the excavation. The article presents an alternative for steel bearing plates, namely plates made using the spatial 3D printing technology. Prototype bearing plates were printed on a 3D printer Formiga P100 using the “Precymit” material. The used printing technology was SLS (Selective Laser Sintering), which is one of the most widely used technologies among all the methods of 3D printing for the short series production of the technical parts of the final product. The article presents the stress–strain characteristic of the long expansion connected rock bolt support OB25 with a length of 3.65 m. A rock bolt support longer than 2.6 m is an additional bolt support in excavations, and it is increasingly frequently used to reinforce roofs and in rebuilding the underground mines of KGHM Polish Copper S.A. In order to conduct the laboratory tests that are most suitable for the mine conditions, and yet are carried out on a laboratory test facility, the Authors used a steel cylinder with an external diameter of 102 mm and a length of 600 mm, which was filled with a core of rock (dolomite) from the roofs of the mine workings. In addition the maximum load that took over the bolt support made of rods and connected with sleeves was determined. For the initial tension, the elastic and plastic range of the maximal displacements, which were measured by the rope encoder, were determined. The statical tests of the expansion rock bolt support were carried out at the laboratory of the Department of Underground Mining in simulated mine conditions. The test facility enables the study of the long bolt rods on a geometric scale of 1:1 for the different ways of fixing. The aim of the laboratory research was to obtain the stress–strain characteristics, of the long expansion rock bolt support with a steel bearing plate and a plate printed on a 3D printer.

The basic type of rock mass reinforcement method for both preparatory and operational workings in underground metal ore mines, both in Poland and in different countries across the world, is the expansion shell or adhesive-bonded rock bolt. The article discusses results of static loading test of the expansion shell rock bolts equipped with originally developed deformable component. This component consists of two profiled rock bolt washers, two disk springs, and three guide bars. The disk spring and disk washer material differs in stiffness. The construction materials ensure that at first the springs under loading are partially compressed, and then the rock bolt washer is plastically deformed. The rock bolts tested were installed in blocks simulating a rock mass with rock compressive strength of 80 MPa. The rock bolt was loaded statically until its ultimate loading capacity was exceeded. The study presents the results obtained under laboratory conditions in the test rig allowing testing of the rock bolts at their natural size, as used in underground metal ore mines. The stress-strain/displacement characteristics of the expansion shell rock bolt with the deformable component were determined experimentally. The relationships between the geometric parameters and specific strains or displacements of the bolt rod were described, and the percentage contribution of those values in total displacements, resulting from the deformation of rock bolt support components (washer, thread) and the expansion shell head displacements, were estimated. The stiffness of the yielded and stiff bolts was empirically determined, including stiffness parameters of every individual part (deformable component, steel rod). There were two phases of displacement observed during the static tension of the rock bolt which differed in their intensity.

The article deals with the solution of analytical problem of flat axisymmetric elasticity theory, when rock-bolt supporting strength is exhausted. Expression for determination of stresses in concrete lining is deduced, where pressure increasing at exploitation have regarded. This methodology can be used at blueprint stage and preliminary calculation during mining operations. Further evaluation of concrete lining strength and geophysical probe of rock is needed for the preliminary calculation.

The position determination for rock bolt support design based on rock stress distribution is important in ensuring full mobilization of rock bolt tensions during excavation process. The monitoring of the progressive rock stress variation in magnitude and orientation can play an important role in optimal installation of rock bolt. The aim of this paper is to determine the optimal positions for rock bolt installation by computing the progressive directional variation of rock stresses and plot the stress distribution in a two-dimensional stereographic projection to capture the stress transformation during excavation process. Finally, an algorithm based on stress transformation determination for optimal rock bolt installation is proposed.

The results of the physical modeling of the “rock bolt – rock massif” system are presented. The characteristics of rock bolt free oscillations depending on the degree of its clamping and tension were studied on special test benches. Oscillations were caused by a single hit to the rock bolt end. It has been established, that the relaxation time of rock bolt free oscillations is the most informative parameter for non-destructive testing. Experimental dependences, that connect the informative parameter with the characteristics of the rock bolt fastening quality, have been determined. An optimal scheme for registering the oscillations in the “rock bolt – rock massif” system is proposed. The basic data has been obtained to substantiate the method of non-destructive testing the quality of rock bolt fastening.

Rock bolting is a method used for rock reinforcement in underground mining and tunnelling. There is a large variety of different types of rock bolts with different support functions. The behaviour of a rock bolt in a rock mass depends on the function and material of the bolt itself, combined with the mechanical properties of the rock mass, deformation capacity, strength and level of stress. Testing of rock bolts in full-scale laboratory-controlled conditions is therefore of great importance. At the rock mechanics laboratory of SINTEF and Norwegian University of Science and Technology (NTNU) in Trondheim, a rock bolt test rig has been developed for full-scale testing for pull, shear and combination pull-shear tests. This paper describes the principles behind this quasi-static full-scale testing and includes the results and analyses of recent tests performed on different types of rock bolts. It also discusses the applicability of the test rig for rock bolt selection and rock support design.

The role of rock bolts has become more important for a tunnel/cavern construction because of the wide use of conventional tunnelling methods under severe conditions. However, there is a lack of understanding on the interface behaviour between a rock bolt and bond material, especially on the crack initiation and propagation. The purpose of this paper is to investigate the effect of bolt configurations on the bolt-grout interface behaviour by discontinuous deformation analysis. By assessing the impact of key parameters, we can better understand the recommended parameter values and their impacts on the failure mechanism.

The article presents methods of making the rock bolt support more yieldable, especially for a stratified roof. Alongside the increasing depth of exploitation of raw material deposits, rock bolt support units are more often designed taking into account more intensive deformations and displacements of underground excavations. In the article, a room and pillar method with mined roof bending and roof reinforcement with bolt patterns of 1 m &times, 1 m, 1.5 m &times, 1.5 m and 2 m &times, 2 m is presented. Moreover, the laboratory tests included 1.8 m long bolts, which were embedded segmentally on the lengths of 100 mm, 150 mm and 200 mm were tested. Based on the load&ndash, displacement characteristics, the deformation energy for flat and profiled dome bearing plates was calculated. Making the segmentally embedded resin rock bolt support yieldable enabled it to perform additional work. Furthermore, it was found that rock bolt support with a dome bearing plate took over 2.5 times more energy compared to a rock bolt support equipped with a flat bearing plate.

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.

Conventional tunnelling methods (i.e., the drill and blast tunnelling method or the new Austrian tunnelling method) have been frequently applied for the construction of tunnels/caverns under a high overburden and with a large cross section. Under such adverse conditions, tunnel support materials may yield because of ground pressure and tunnel deformation. As a result, a tunnel may lose its stability due to a reduced effectiveness of the tunnel supports. Rock bolts have been widely used as an essential component of tunnel support, and many studies have been conducted on the performance of rock bolts in strengthening the jointed rock mass. However, there is a lack of understanding on the interface behaviour between the rock bolt and the bond material, especially the crack initiation and propagation inside the bond material. This study aims to investigate the interface behaviour between rock bolts and bond materials using laboratory tests (shear tests) and numerical simulations (Discontinuous deformation analysis). By assessing the impact of the key parameters in the rock bolting system, such as the ribs, rib angle, strength of the bond material and the confining pressure, we are able to better understand the supporting mechanism and the effects of rock bolting.

A series of laboratory tests were performed to study the mechanical behaviours of newly developed high strength rock bolt components, including rebar, thread, plate, and domed washer. The characteristics of deformation and damage of each component were presented. The stress distribution of plate and domed washer was investigated through finite element modelling. The numerical results show that the yield and tensile strengths of the developed high strength rebar are 33.6%–58.3% and 17.2%–28.7% greater than those of the conventional rebar, respectively. The increase in yield strength was higher than that in tensile strength, suggesting an increase in yield to tensile strength ratio and a decrease in elongation. It is well-known that the thread processing may not be of high precision and accuracy as expected, which is characterised as rough thread surface, non-identical tooth height, toe stripping, and cracks in the surface. Hardening during thread processing tends to increase the thread yield and tensile strengths. In this paper, the typical deformation process of arch-shaped plate is classified into five stages. The tested plates exhibited distinct deformation characteristics and bearing capacities due to variations in shape, size, material and presence of washer. It was observed that uneven bottom surface, low bearing arch and large radius of the transitional arc connecting bearing arch and bottom surface were the major reasons accounting for low load-bearing capacity of plates. The performance of domed washer has a close relation with the shape, size, strength, and deformation compatibility with plate. Stress concentration was observed on the periphery of the contact surface between domed washer and plate, which is significantly influenced by the strength of domed washer and is considered to be 20%–30% higher than that of plate. Finally, a case study in the Datong coal mining district was presented, and the support pattern and effect of the developed rock bolt were described.

Given the increase in mining depth and intensity, tunnel failure as a result of rock burst has become an important issue in the field of mining engineering in China. Based on the Composite Rock-Bolt Bearing Structure, which is formed due to the interaction of the bolts driven into the surrounding rock, this paper analyzes a rock burst prevention mechanism, establishes a mechanical model in burst-prone ground, deduces the strength calculation formula of the Composite Rock-Bolt Bearing Structure in burst-prone ground, and confirms the rock burst prevention criterion of the Composite Rock-Bolt Bearing Structure. According to the rock burst prevention criterion, the amount of the influence on rock burst prevention ability from the surrounding rock parameters and bolt support parameters is discussed.

In this paper based on the explosive stress wave theory, the mechanism for the broken rock zone’s formation and the character of the surrounding rock’s disfeature under blast loads are analyzed. On the basis of these works, the surround rock’s dynamic broken rock zone calculation formula is derived and for the anti-explosion reinforcement to the surrounding rock of tunnel segment bearing dynamic load is locked. In addition, the author also calculated and analyzed separately the necessary length of rock bolt for the reinforcement, consider of the first type to the forth type surrounding rock, different vertical critical explosion distance, different span. The curves are obtained between the length of rock bolt for the anti-explosion reinforcement of the surrounding rock and three factors are included: the span of underground opening, type of the surrounding rock, the distance between the center of charge and the vault.

The purpose of this study was to investigate the effect of bolt profile on load transfer mechanism of fully grouted bolts in jointed rocks using analytical and numerical methods. Based on the analytical method with development of methods, a new model is presented. To validate the analytical model, five different profiles modeled by ANSYS software. The profile of rock bolts T3 and T4 with load transfer capacity, respectively 180 and 195 kN in the jointed rocks was selected as the optimum profiles. Finally, the selected profiles were examined in Tabas Coal Mine. FLAC analysis indicates that patterns 6+7 with 2NO flexi bolt 4 m better than other patterns within the faulted zone. Keywords: Fully grouted bolt, Load transfer mechanism, Jointed rocks, Analytical and numerical methods

The effect of long term exposure of full size bolts to corrosive environments is presented. A special test rig was used to test four bolts under different loading conditions. Four X-grade identical profile bolts, each 21·7 mm core diameter (23·7 mm full diameter) were subjected to prolonged corrosion testing using acid sulphate water. The pH value of the circulated water varied between 3·4 and 4·3. The corrosion exposure test period lasted 3·5 years. Two bolts were axially loaded to 10 t and 20 t force respectively, the third bolt was subjected to a 360 Nm torsion load and the fourth bolt was left unstressed to act as a reference bolt. After the test period ended, the bolts were stripped of their corroded coatings and weighted for weight loss. The diameter of each bolt was subsequently measured, and the loaded bolt samples were first tested non-destructively for tensile cracks and then tested for tensile failure. No cracks were found on post-corrosion bolts tested non-destructively. The failure st...

Analysis of mechanical behaviour of rock mass reinforced by fully grouted rock bolt is introduced based on the enriched finite element method (EFEM). A solid element intersected by a rock bolt along any arbitrary direction is defined as ‘enriched’ element and it has additional degrees of freedom at each node for estimating displacements and stresses in bolt rod. Numerical procedures of EFEM are developed to form enriched stiffness matrix using constitutive relations of rock mass, properties of bolt rod and grout material and orientation of the bolt. Decoupling of rock bolt and elasto-plastic behaviour of rock mass has been incorporated into the EFEM procedures. Released displacement of rock mass prior to bolt installation has also been considered in the procedures to provide realistic solution. The results of this method are verified with an analytical pull-out test model. In addition, a numerical example of a bolted tunnel is provided to demonstrate the efficacy of the proposed method for practical applications.

This study analyzes the phenomenon of performance deterioration in fully grouted rock bolts in tunnels with a dry, hot environment and high geothermal activity with a focus on temperature effects on interfacial bond performance. Three groups of fully grouted rock bolt specimens were designed based on similar mechanical principles. They were produced and maintained at 20 °C, 35 °C, and 50 °C. Through the indoor gradual loading tensile test of specimens, variations of axial force and shear stress between the rock bolt and mortar adhesive interface were obtained under different environmental temperatures. Distribution of the axial force and shear stress on the anchorage section were found under different tensile forces. Results showed that, with an increase in specimen environmental temperature, maximum shear stress of the rock bolt section became smaller, while shear stress distribution along the rock bolt segment became more uniform. In addition, the axial force value at the same position along the pull end was greater, while axial stress along the anchorage’s length decayed faster. With an increase in tensile force under different temperatures, the axial force and maximum shear stress of rock bolt specimens along the anchorage section has a corresponding increase.

An analytical model for rock bolts has been developed based on an improved Shear–Lag Model. The development of the model is based on the description of the interaction behavior of the rock bolt, the grout medium and the rock mass. On the basis of the model, the coupling and decoupling behavior of the rock bolt in pullout tests, uniform deformation the rock mass and intersecting joints are analyzed. The pullout test characteristic is described by the proposed model, and a back analysis method is proposed to calculate the shear strength of the interface media. For the rock bolts in a deformed rock mass, the influence of the installation time of the rock bolt has been taken into account, and the theoretical prediction is verified by the measured data. According to the proposed model, the position of the neutral point is not only related to the length of the rock bolt and the radius of a tunnel, but also is strongly influenced by the mechanical properties of the rock mass. Analysis of the joints intersecting the rock bolt shows that there may be more than one neutral point on the rock bolt, and the prediction of the simplified model is consistent with the pullout model. By using this model, a method is proposed to analyze the interaction behavior of the rock bolt and the surrounding rock mass, with a way of evaluating the supporting performance quantitatively.

A research programme into both the static and dynamic performance of ground anchorage systems, with an emphasis on resin bonded rock bolts, started at the University of Aberdeen in the early 1980s. The work involved measurements on active construction sites which was underpinned by laboratory and computer modelling and led to the development of a new method for the non-destructive testing of anchorages. Part of the research programme was focused on assessing how changes in the load influence the dynamic responses of an anchorage. This in turn produced a new testing method able to determine the variation in frequency response of the anchorage with changes in load. The development of a lumped parameter model able to simulate the response of anchorages to changes in static load and an applied impulse load was a further step forward in the research programme. The achievements of the lumped parameter model revealed the head of the anchorage to be the most influential component of the anchorage system in determining dynamic response. The results obtained from the numerical model when laboratory anchorages were simulated are shown to be in agreement with the results obtained from the actual laboratory tests. Based on the successful results obtained to date regarding laboratory rock bolt anchorages, a further step was made in order to employ the numerical model for the first time to observe the influence of load on the frequency response of rock bolt anchorages installed in the field. This paper describes the application of the model to such field anchorages and highlights some of the detailed modelling aspects required to replicate real anchorage behaviour.

鋼管膨張型ロックボルトは湧水のある地山や掘削による初期変位のあるトンネル工事で使用される。湧水中での耐久性を評価するために，溶融Zn-6％Al-3％Mg合金めっき製ボルトおよびめっきなしボルトを東海北陸自動車道飛騨トンネルの湧水地山に試験設置した。1，2，3年経過後にオーバーコアリング法によりボルトを地山から回収し，腐食挙動を調査した。めっきなしボルトの場合，腐食（板厚減少）速度は直線近似でき，溶存酸素拡散支配型の腐食と判断された。一方，合金めっき製ボルトの場合，灰色外観を呈しほとんど腐食の認められない領域とめっき層が局部的に腐食した黒変色領域とが混在していたが，黒変色領域における腐食の進行も2年経過以降はほぼ停止していると見なされた。3年経過後の灰色領域，黒変色領域および溶射補修した溶接部のいずれの部位も最表層はZn-Si-O系の均一な皮膜で覆われており，試験設置場所で採取した湧水の分析結果などからケイ酸亜鉛皮膜と推定された。このケイ酸亜鉛皮膜の腐食抑制効果により，それ以上の腐食の進行がほぼ停止した状態になっていると考えられる。

2.4m立方の模型地山に60cmのトンネルを設け, ロックボルトパターンを種々変えて載荷試験を行い, ボルトのトンネル補強効果と最適パターンの研究を行った. 載荷は側圧係数1.0と0.5の二通りで行った. 模型地山には一軸強度1.0MPa, 内部摩擦角30の砂質低強度モルタルを用いた. ボルトは長さ40～10cm, 本数は8cm間隔の1断面に12～48本とした. この実験の範囲では, 短いボルトを数多く用いることの有効性が明らかとなった.

Yielding rock bolt support systems have been developed to accommodate ground movement in shifting ground such as in coal operations; in creeping ground such as salt, trona, and potash; and in swelling ground associated with some clays. These systems, designed to remain intact despite ground movement, should enhance mine safety and help contain costs in areas where revolting of rigid non-yielding systems is typically required. Four such systems were tested in straight tensile pulls in the laboratory. They include the Slip Nut System from Dywidag Systems International USA, Inc., Ischebeck`s bolt mounted Titan Load Indicator, Rocky Mountain Bolt Company`s Yielding Cable Bolt, and a rock bolt installed variation of the yielding steel post developed by RE/SPEC Inc. The first two systems are currently marketed products and the latter two are prototype systems. Each system responds to load and displacement by yielding in a unique manner. All are designed to yield at predetermined loads. A description of each system and its yield function is provided. Each system was tested over its prescribed yield range in a test machine. At least five tests were performed on each system. Each system yielded and continued to provide support according to its design. Each showsmore » promise for ground control use in shifting or creeping rock. This work helps to illustrate the comparative differences in performance between these specialized systems and the applications where they may be most useful.« less

Rock mass classification has played a crucial role in underground construction and mining projects in the past fifty years, especially the Rock Mass Rating (RMR) and Rock Tunneling Quality Index (Q) systems, which have been applied in many cases. The parameters of discontinuity conditions in the RMR system, along with the related ratings, were used to develop a new equation in order to introduce an intrinsic quality of a given rock mass, namely, rock bolt supporting factor (RSF). The RSF was used to develop a mathematical theory of the rock bolt supporting mechanism as a new principle in explaining rock bolting effects. Finally, using the parameters of the discontinuities in the Q-system, the QRSF is defined to present the capability of a given rock mass to be reinforced by rock bolting. The RSF leads to more reliable judgment of rock bolting capability of a given rock mass than the QRSF since it uses five parameters of discontinuity conditions while the QRSF uses only two parameters.

Despite their frequent natural occurrence and engineering encounter, heterogeneous rock joints (rock joints with different lithological characters on both sides of the joint surface) have been studied much less systematically. To study the shear behavior and failure mechanism of bolted heterogeneous rock joints, laboratory tests were performed on the heterogeneous rock joints having different joint roughness coefficients (JRC) under different anchorage conditions. The results indicate that shear strength increases with the increase of JRC, showing exponential growth. Under the same roughness, the shear strengths of rock joints from large to small are fully grouted, end anchorage, and without anchorage. The mechanical characteristics of the bolt and joint are poorly matched under the end anchorage condition, which is easy to cause these two to be broken one by one. Under fully grouted, the extrusion force caused by the rock bolt will diffuse around the anchorage agent and will not cause partial continuous damage. The surface damage of heterogeneous rock joints increases with the increase of JRC and presents obvious heterogeneous characteristics. The shear dislocation between the blocks under shear load results in the interaction between the bolt and surrounding media. Under the action of shear force, the bolt body produced both axial and transverse deformation, which leads to breakage of anchorage agent and rock mass. Rock bolt has a significant impact on the shear behavior of the anchorage system, and the damage of the rock bolt to rock mass should be considered in rock engineering reinforcement design.

Understanding the mechanism of progressive debonding of bolts is of great significance for underground safety. In this paper, both laboratory experiment and numerical simulation of the pull-out tests were performed. The experimental pull-out test specimens were prepared using cement mortar material, and a relationship between the pull-out strength of the bolt and the uniaxial compressive strength (UCS) of cement mortar material specimen was established. The locations of crack developed in the pull-out process were identified using the acoustic emission (AE) technique. The pull-out test was reproduced using 2D Particle Flow Code (PFC2D) with calibrated parameters. The experimental results show that the axial displacement of the cement mortar material at the peak load during the test was approximately 5 mm for cement-based grout of all strength. In contrast, the peak load of the bolt increased with the UCS of the confining medium. Under peak load, cracks propagated to less than one half of the anchorage length, indicating a lag between crack propagation and axial bolt load transmission. The simulation results show that the dilatation between the bolt and the rock induced cracks and extended the force field along the anchorage direction; and, it was identified as the major contributing factor for the pull-out failure of rock bolt.

Numerical simulation is a useful tool in investigating the loading performance of rock bolts. The cable structural elements (cableSELs) in FLAC3D are commonly adopted to simulate rock bolts to solve geotechnical issues. In this study, the bonding performance of the interface between the rock bolt and the grout material was simulated with a two-stage shearing coupling model. Furthermore, the FISH language was used to incorporate this two-stage shear coupling model into FLAC3D to modify the current cableSELs. Comparison was performed between numerical and experimental results to confirm that the numerical approach can properly simulate the loading performance of rock bolts. Based on the modified cableSELs, the influence of the bolt diameter on the performance of rock bolts and the shear stress propagation along the interface between the bolt and the grout were studied. The simulation results indicated that the load transfer capacity of rock bolts rose with the rock bolt diameter apparently. With the bolt diameter increasing, the performance of the rock bolting system was likely to change from the ductile behaviour to the brittle behaviour. Moreover, after the rock bolt was loaded, the position where the maximum shear stress occurred was variable. Specifically, with the continuous loading, it shifted from the rock bolt loaded end to the other end.