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Rock bolts are commonly used to stabilise slopes and underground excavations but their installation can sometimes be difficult in challenging grounds due to excessive leakage or washout of the grout. To manage these problems, tubular geotextile socks have been adopted in the industry but there has been a lack of research on their properties. To advance the understanding of sock behaviour, the water permittivity and filtration properties of two types of grout sock have been investigated in their unstretched state. From the permittivity tests, it is shown that water flow through these socks will be laminar and that using more than one layer in a setup can effectively reduce the flow velocity. Moreover, it has been shown that the effect of the tester should be considered in the analysis or else the permittivity of the geotextile could be underestimated. From the filtration tests, it is shown that both types of grout sock tested are effective in minimising grout loss and that the permittivity of the filter cake will be many orders of magnitude lower than that of the socks. The results also show that the water–cement ratio of the grout will be reduced as a result of the dewatering process.

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.

Rock bolt anchorage performance is crucial for tunnel support safety. We investigate the mechanical behaviour of reinforced rock and the bolts that reinforce it from the micro-scale to the macro-scale. Bolt pull-out tests were performed on soft rock using the distinct element method, in which a new contact model that considers bond size, is employed to constrain the main rock mechanical behaviour. The minimum sample width and height values for which the boundary effect can be neglected are first proposed through numerous tests on the influence of sample size on peak load and bond breakage. The influence of sample width is substantially greater than that of sample height. We then select an appropriate sample size to study the influence of bolt embedment length and confining pressure on the mechanical behaviours of the rock and bolt. The results show that increased rock bolt embedment length and confining pressure can increase the peak load; however, the bolt length effect is limited when exceeding the critical anchorage length. In cases without confining pressure, bond breakage occurs in the rock around the grout-rock interface and the breakage zone is rectangular, whereas in cases under confining pressure, the breakage zone presents an inverted cone shape. We use our results to discuss the influence of bond strength at the bolt–grout interface on the peak load and failure mode. The failure mode changes gradually from complex failure to single failure along the bolt–grout interface with decreasing interfacial bond strength.

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.

This paper presents results of 71 in situ nondestructive rock bolt and tendon pull tests conducted in the Powerhouse Complex of the Uma Oya Multipurpose Development Project, located in Sri Lanka. The pull tests are performed on 6 and 8 m Double Corrosion Protected fully and partially grouted rock bolts (high strength threadbar) as well as 4 m normal fully grouted rock bolts and 26 m pre-stressed tendons, which are installed in different types of Gneiss. Based on the recorded load–displacement curve of the rock bolt head in each test, stiffness of fully and partially grouted rock bolt is measured in the bonded section. Effect of some parameters on the stiffness such as rock mass types and characteristics, bond length as well as rock bolt diameter and steel type are investigated. Primarily, the pull tests are done to approve the working load capacity of rock bolts and tendons and quality control of the installation process. Besides, the stiffness of bonded rock bolt is also calculated. The results show that the stiffness declines by decreasing the rock mass mechanical characteristics, rock bolt bond length, and rock bolt diameter as well as steel bar grade.

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.

The corrosion of rock bolts buried in rock and soil layer is obviously different from that of general ground structure. The main corrosion sources of the latter come from the air and the corrosion sources of rock bolts are not only from the air, but also the rock-soil body and groundwater. At the same time, the corrosion of rock bolts is under the stress state in practical engineering. To study the stress-corrosion effect of grouted rock bolts, using a self-developed rock bolt stress corrosion device, corrosion tests were conducted under different working stresses (10 kN, 35 kN and 60 kN), corrosion environments (5% Na2SO4, 1.5% NaCl + 2.5% Na2SO4 and 3% NaCl) and corrosion times (3 d, 7 d and 14 d). The results show that the NaCl solution is more corrosive than the Na2SO4 solution for the grouted rock bolts with more rust productions. The maximum tension of rock bolts under the condition of no corrosion reaches 60.16 kN, and under the T-60-A-14d working condition, the maximum tension of rock bolts is only 31.32 kN. Compared with the condition of no corrosion, the maximum tension of grouted bolts is lowered by 25%–50% under the different working stresses and corrosion conditions. The maximum difference values of the relative bonding strength of rock bolts is 14.9% under corrosion time, 13.1% under working stress and 9.4% under corrosion environment. It shows that the corrosion effect of the rock bolt under the corrosion time is the most significant, and the working stress is second. The influence of the corrosion environment on the bonding strength of the grouted rock bolts is the least important. The slip curves of the rock bolts under the environment of no corrosion have an obvious rising stage, and uplift of tension accounts for about 62.9% of the maximum tension. The final residual tension is stable at approximately 25 kN. The bonding slip curves of corroded rock bolts, of which the “initial tension rising stage” almost disappears, demonstrate a trend of steady decline. The minimum final residual tension is less than 10 kN (T-35-B-14d). The larger the working stress is, the more significant the degradation effect of anchor tension is. It is shown that the current test and study on the corrosion of anchor bolts can not accurately reflect the corrosion law of rock bolts in practical engineering if only the material or the axial force with small value is applied. The influence of stress on the corrosion test of rock bolts can not be ignored.

High stress in surrounding rock mass could cause serious stability problems such as larger squeezing deformation in soft rock and rock burst in hard rock. The support system applied in high in situ stress conditions should be able to carry high load and accommodate large deformation of rock mass. This paper presented a new yielding rock bolt, called tension and compression-coupled yielding rock bolt, which is promising to provide support for both squeezing and burst-prone rock mass encountered in mining or tunneling at depth. The new bolt mainly consists of a steel rod and two additional anchors. The steel rod is a round shape bar with varying surface conditions. The inner segment is processed into rough surface, while the middle segment of the rod has smooth surface. Two additional anchors were arranged on both ends of smooth segment. The bolt is fully encapsulated in a borehole with either cement or resin grout. The rough segment and the inner anchor are firmly fixed in the bottom of the borehole, while the smooth segment has no or very weak bonding to the grout, which can stretch to accommodate rock dilatation. First, direct quasi-static pull tests were performed to examine the load capacity of tension and compression-coupled anchor. The results showed that the coupling action of tension to the rough rod and compression on the inner additional anchor by grout in different positions can increase the ultimate bearing capacity of inner anchoring segment significantly. Second, the performance of the new bolt and the fully encapsulated rebar bolt was tested under fracture opening condition. Results showed that the load and strain concentration could result in premature failure of fully encapsulated rebar bolt. However, the smooth segment of TCC Yielding rock bolt can detach from the grout under pull loading and provide a larger deformation to accommodate rock dilations. Third, shear tests were performed to examine the deformation mechanism of the new rock bolt under fracture sliding condition. Results showed that the smooth section of the new bolt specimen can deform freely to accommodate the sliding of fracture. The maximum shear displacement of the new bolt specimen is much larger than the fully encapsulated rebar bolt specimen, which is promising a better ability to accommodate the large displacement sliding of fracture in engineering practice.

The shear performance of bolted rock joints under cyclic loading condition was studied through laboratory shear tests. The tests were performed on a cyclic shear testing apparatus. The fully encapsulated rock bolts were inserted in the rock joints. The upper block was fixed in the horizontal direction, and the cyclic shear load was applied laterally on the lower block. A total of 16 shear tests were conducted for two types of rough joints. The results showed that the failure mode of rock bolts in cyclic shear test was different with the case of direct shear test. When the cyclic displacement was small, the rock bolt played a very small role in the second-to-fifth cycles, the rock bolt remained unbroken during cyclic shearing. The shear resistance of the rock bolt would gradually recover after the shear displacement has exceeded the cyclic displacement in the termination cyclic after five cycles. When the cyclic displacement was large, the rock bolt broke during cyclic shearing, and the shear resistance could not be recovered in the termination cyclic. It is indicated that the rock bolt had completely lost its supporting role after cyclic shear loading. The shear strength reduction of bolted joints under the cyclic loading condition was much significant compared to the un-bolted joints. Therefore, the influence of cyclic loading on the shear resistance of the rock bolt was much larger than its counterpart of the rock joint itself. These results showed that the shear performance of a rock bolt inserted in a rock joint was strongly influenced by cyclic shear loading.

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.

This paper aims to propose new numerical procedures for elasto-plastic analysis of the interactions between fully-grouted bolt and rock with slip and non-slip cases at a potential-failure interface. In the proposed procedure with the slip case, the iteration process accounts for the elongation of the bolt, and the calculation process is greatly simplified by considering only the initial and final states (i.e., no intermediate finite states as is the case for existing procedures) without compromising the accuracy of the predictions. In the proposed procedure with the non-slip case, the elasto-plastic solution of the plastic rock strain is rigorously derived; the two scenarios in which the length of the rock bolt is contained within the plastic zone or within the elastic zone are comprehensively considered. The obtained solutions of tensile force in the bolt, shear stress at the potential-failure interface, rock displacement and stress components in the rock from the proposed procedures are validated by analyses with purely analytical methods and a commercial finite-difference program. With influences of rock quality and plastic softening behaviour of rock mass, the differences of the solutions in the slip and non-slip cases at potential-failure surface are compared; the dependency of the relative movement on the shearing stiffness at the potential failure interface is also investigated.

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.

Rock bolts are commonly used in conjunction with Fibre Reinforced Shotcrete (FRS) linings to stabilise ground around excavations. A steel cover plate is often placed over the protruding end of a rock bolt to dissipate load from the bolt into the FRS lining. The process of drilling and installing rock bolts takes place after the lining is sprayed, so a hole must be made through the lining to insert the bolt. To examine the effect of this on lining behaviour, an investigation has been carried out to determine the influence of a central hole on the punching capacity of a FRC panel. Grout or concrete is commonly placed between the cover plate and lining to distribute the load more evenly, so the effect of this aspect of rock bolting practice on punching behaviour has also been examined.

The existence of uncertainty in the design parameters of anchored rock slopes warrants reliability based study. An approach for reliability based design optimization of anchored rock slopes against plane failure is described under different hydraulic distributions. Planar slide failure along with water pressure has been considered in conjunction with Barton-Bandis failure criterion to compute the factor of safety. Nonlinear strength criterion is employed for the computation of strength of jointed rock mass. The rock bolt and rock mass interaction model is used to analyze the stabilizing force in anchors. The variability associated with Barton-Bandis parameters, namely joint compressive strength (JCS), joint roughness coefficient (JRC), and basic friction angle (ψ) and a joint factor is considered. The target reliability based design optimization (TRBDO) is used to obtain an optimum length of the rock bolt and optimum anchor force required to stabilize the rock slope in order that the reliability index must meet a desired level of safety.

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.

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.

Rock bolts are commonly used on transportation projects to stabilize cut slopes, reinforce undermined or unsupported blocks, and reduce potential for large rock slides. Installing bolts can be challenging. Locations where bolts are useful may not be easily accessible for personnel or large equipment, and targeted blocks generally are marginally stable. In addition, jointed rock with openings and voids, which are typical conditions, complicate hole advancement and full grout encapsulation. These issues can lead to construction delays, claims, reduced corrosion protection, or shorter design life. This paper describes research conducted to address two rock bolting challenges in open-jointed rock: i) advancement of drill holes to design depth, and ii) successful completion of grouting through the application of injected polyurethane foam (PUF). Testing utilized double-stacked concrete blocks aligned in rows to simulate a rock mass with vertical fractures. Steel spacers were placed between the stacks to simulate 2-, 4-, 6-, and 8-inch apertures. A standard rock drill advanced holes using a 3-inch bit. Research showed PUF could: bridge simulated aperture sizes up to 8 inches, survive drilling action during secondary reaming, resist drilling air and grouting pressures, and reduce grout consumption.

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.

In underground coal mining, high abutment loads caused by the extraction of coal can be a major contributor to many rock mechanic issues. In this paper, a large-scale physical modeling of a 2.6 × 2.0 × 1.0 m entry roof has been conducted to investigate the fundamentals of the fracture mechanics of entry roof strata subjected to high abutment loads. Two different types of roof, massive roof and laminated roof, are considered. Rock bolt system has been taken into consideration. A distinct element analyses based on the physical modeling conditions have been performed, and the results are compared with the physical results. The physical and numerical models suggest that under the condition of high abutment loads, the massive roof and the laminated roof fail in a similar pattern which is characterized as vertical tensile fracturing in the middle of the roof and inclined shear fracturing initiated at the roof and rib intersections and propagated deeper into the roof. Both the massive roof and the laminated roof collapse in a shear sliding mode shortly after shear fractures are observed from the roof surface. It is found that shear sliding is a combination of tensile cracking of intact rock and sliding on bedding planes and cross joints. Shear sliding occurs when the abutment load is much less than the compressive strength of roof.

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.

This paper presented solutions of displacement and stress for a circular opening which is reinforced with grouted rock bolt. It satisfies the Mohr-Coulomb (M-C) or generalized Hoek-Brown (H-B) failure criterion, and exhibits elastic-brittle-plastic or strain-softening behavior. The numerical stepwise produce for strain-softening rock mass reinforced with grouted rock bolt was developed with non-associative flow rules and two segments piecewise linear functions related to a principle strain-dependent plastic parameter, to model the transition from peak to residual strength. Three models of the interaction mechanism between grouted rock bolt and surrounding rock proposed by Fahimifar and Soroush (2005) were adopted. Based on the axial symmetrical plane strain assumption, the theoretical solution of the displacement and stress were proposed for a circular tunnel excavated in elastic-brittle-plastic and strain-softening rock mass compatible with M-C or generalized H-B failure criterion, which is reinforced with grouted rock bolt. It showed that Fahimifar and Soroush's (2005) solution is a special case of the proposed solution for n = 0.5. Further, the proposed method is validated through example comparison calculated by MATLAB programming. Meanwhile, some particular examples for M-C or generalized H-B failure criterion have been conducted, and parametric studies were carried out to highlight the influence of different parameters (e.g., the very good, average and very poor rock mass). The results showed that, stress field in plastic region of surrounding rock with considering the supporting effectiveness of the grouted rock bolt is more than that without considering the effectiveness of the grouted rock bolt, and the convergence and plastic radius are reduced.

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.

Understanding the interaction between rock bolts and underground rock movement is critical for safe and cost-effective underground excavation design. Although early research on this subject involved a balance of theoretical analysis and field measurement, the current work focused on analytical and numerical studies. The present study aimed to study fully grouted rock bolt in Tabas Coal Mine using numerical and instrumentation methods. Tabas Coal Mine is the first fully mechanized coal mine in Iran, producing 1.5 million tons of coal per year. The mine extracts coal by both longwall and room and pillar methods. The roadways have a rectangular profile of 4.5 m width and 3.5 m height. The field investigations and geotechnical characteristics of rocks showed that the rock masses are weak, requiring a suitable support system. The roadway is intersected by a major fault zone. For the designed roadway, four patterns (two types of bolt with transfer load capacity of 180 and 195 kN) were considered. FLAC analysis showed that 13 bolt 2.4 m with 2 NO flexi bolt 4-m patterns were better than other patterns within the faulted zone. Assessing the optimum pattern through instrumentation, there was compromise between the results of numerical modeling with instrumentation reading such as telltales, sonic extensometer, and strain gauge rock bolt. Thus, there was an inspiring agreement between numerical and field methods.

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.

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.

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.

In this study, usability of Polyamide-6 type engineering polymer as a new rock bolt material was investigated carrying out a series of laboratory and field scale experiments. The Polyamide-6 type polymer tested as a grouted rock bolt material was assessed to have a good body strength, large elastic deformation limit and high energy absorbing capacity. The load bearing capacity of Polyamide-6 rock bolts was found to dominantly depend on the mechanical anchorage in front of the shank as a result of having low adhesion to the cement grout values. In this study, polyamide rock bolts have been designed with different mechanical anchors. It was suggested to start the systematical use of polyamide rock bolts having ideal support reactions under both static and dynamic load conditions in rock engineering. Especially, Polyamide-6 rock bolts were assessed to be economically usable in rock masses with squeezing, swelling or bursting problem, which need for non-decreasing support pressure while being higly deformed and good energy absorption capacity.