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Slope instability is a complex geological phenomenon triggered by heavy rainfall, earthquakes, tectonic forces and anthropogenic activities. Unplanned excavation of rock slopes for the development and maintenance of infrastructure such as highways, railways, and buildings in hilly regions plays a significant role in causing slope instability. The present work conducts a comprehensive stability assessment of rock slopes along the national highway (NH-44) section of the Ramban district of Jammu and Kashmir. The kinematic investigation was carried out to determine the different modes of failure of road-cut slopes, and two main modes were found: planar and wedge. Various empirical classifications have been applied for the assessment of slope stability, like geological strength index (GSI), slope mass rating (SMR), Chinese slope mass rating (CSMR), and continuous slope mass rating (CoSMR). Numerical analysis through universal distinct element code (UDEC) has been conducted to evaluate the stability of four critical rock slopes under static and dynamic loading conditions. Modelling results have provided insights into the failure mechanism, and based on these findings, remedial measures have been proposed. Three (L-1, L-3, and L-13) out of four slopes were determined to be unstable under static conditions, with factors of safety (FoS) < 1.2 and under dynamic conditions, two slopes were unstable with FoS < 1.0. However, one (L-8) slope was found to be stable under both static and dynamic conditions, with FoS 1.29 and 1.12, respectively. Installing rock bolts reduced displacement for slopes L-1 and L-13, with reductions of 11.86% and 21.05% under static conditions and 4.09% and 18.75% under dynamic conditions, respectively, while L-3 does not stabilise even after installation of rock bolts. [ABSTRACT FROM AUTHOR]

In tunnel construction, grouting is frequently used to decrease the permeability and increase the rock mass strength. One of the most challenging steps in grouting is choosing the best grout type to achieve the highest penetration and sealing efficiency in the joints. The main purpose of this study is to investigate the sealing efficiency of different grout types in the joints of a high-risk section of the Kerman water conveyance tunnel (KWCT). For this purpose, the penetrations of three types of grouts including Ordinary Portland Cement (OPC), Ultra-Fine Cement (UFC), and silica sol were determined using analytical equations. Thereafter, the sealing efficiency of each type of grout in joints with different apertures was calculated. Also, the penetration radius of each type of grout was simulated and calculated using UDEC code. Based on the results of the analytical equations, by performing the grouting for 20 min using silica sol-based grout, the joints of the high-risk section of the KWCT with apertures of 0.5–2.5 mm and 2.5–10 mm can be sealed at 'Fair' and 'Excellent' levels, respectively. Furthermore, based on the results of numerical simulations, the UFC and silica sol-based grouts can seal the joints in 'Good' and 'Excellent' levels, respectively. [ABSTRACT FROM AUTHOR]

Abstract The extraction of underground coal resources induces the fracture and movement of overlying strata, leading to geological hazards such as surface deformation, cracks, and even subsidence. Utilizing the analogous hyperbola model of overlying strata movement, we conducted a mechanical analysis to examine the asymmetric fracture mechanism resulting from coal seam mining in thick loose strata. An asymmetric analogous hyperbola model was established by introducing distinct virtual half-axis lengths (b). The thickness impact of thick loose layers (H) and bedrock layer (h) on the asymmetric movement of overlying rock during mining was also discussed. Similarity model tests were conducted to research the migration characteristics and surface subsidence patterns of overburdened rock and thick loose layers at different mining stages and validate the hypothesis of asymmetric overburdened rock migration. Additionally, the discrete element numerical model for thick and loose layers mining was established by using UDEC and discussed the asymmetric analogous hyperbola behaviour of overburden movement and surface subsidence. The comparison results show that the established asymmetric hyperbolic model can effectively predict the movement law of overlying strata and surface subsidence characteristics. Therefore, the proposed model can provide valuable theoretical support for predicting the movement patterns of overburden under thick loose layers and mitigating surface subsidence disasters.

Rock anchors, which can withstand substantial forces and are essential for securing the foundations of the built environment, can fail due to steel tensile failure, bond detachment, or rock mass fracturing through uplift. In this study, uplift failure of the rock mass surrounding rock anchors was simulated using both physical and numerical models. The physical models used a specially designed testing rig with predefined sizes and different joint layouts to simulate rock mass uplift failure. The numerical models were calibrated using the physical models and used to investigate scenarios involving increased complexity and greater depths. This paper presents the load distribution within rock masses with orthogonal joints when exposed to an anchor load. Extensive analysis of the numerical models revealed that symmetrical load arches were induced in rock masses that contained a joint set oriented parallel to the loading direction. However, joints that were obliquely oriented relative to the anchor axis induced asymmetrical load arching. The greatest displacement occurred in the direction parallel to the joint set with the smallest angle to the anchor loading axis. Large-scale models with zero or near-zero in situ horizontal stress revealed that the load capacity in a rock mass with continuous vertical joints was relatively low, especially when compared to rock masses with discontinuous vertical joints. It was observed that the uplift capacity of the rock mass around an anchor decreased if the load arches of adjacent anchors overlapped; consequently, a decrease in anchor spacing was found to correspond to a decrease in uplift capacity. Highlights: Symmetrical load arches are induced within a rock mass if the joint sets are oriented parallel to or close to parallel to the anchor axis. An oblique orientation of the joints relative to the anchor axis results in asymmetrical load arching in the rock mass when the anchor is loaded. Rock masses with continuous vertical joints have a low load capacity, especially compared to rock masses with discontinuous vertical joints. In a row of rock anchors, the rock mass uplift capacity of individual anchors is decreased if the anchor spacing is too low. [ABSTRACT FROM AUTHOR]

Ground vibration is one of the detrimental effects associated with blasting that can damage the surrounding environment and nearby structures. In the Gol-Gohar mine in Sirjan, due to the surface expansion, the distance between the structures and the blasting blocks has decreased, leading to vibrations reaching the processing plant complex. These vibrations, by triggering sensors installed on the mills, cause power outages in the circuit, thereby increasing production costs. One solution to mitigate the waves reaching the processing plant complex is to excavate trenches along the wave path. These trenches, by creating conditions like a free surface and reflecting the waves, reduce the transferred wave energy and can prevent unnecessary shutdowns of the concentration circuit due to increased vibration amplitudes. In this study, using the discrete element software UDEC, the results of a field blasting operation were first validated, and based on the validated model, the impact of trench excavation on the propagation of blast waves was analyzed. Ultimately, the optimal dimensions of the trench, which maximizes energy absorption, were determined. According to the numerical analysis results, the excavated trench on each side of the structure should be more than 2m longer and excavated at distances greater than 3m from the structure. Meanwhile, the thickness (width of the trench) had no significant effect on wave attenuation. This trench can reflect approximately 60% of the blast waves.

Copyright of Transportation Science & Technolgy is the property of Transportation Science & Technology Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Journal of Engineering Geology / Gongcheng Dizhi Xuebao is the property of Journal of Engineering Geology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Abstract The deformation and damage of soft rock roadways in karst mountainous area is serious, which greatly restricts the safe and efficient mining of coal resources. To explore the instability mechanism of soft rock roadway under the influence of mining, the 226 rail cross-cut of Shanjiaoshu Coal Mine in China was taken as an engineering background. The methods of field monitoring and drilling imaging were adopted to analyze the roadway failure characteristics. In addition, the distributions of the plastic zone, surface displacement, and fracture fractal evolution characteristics of the surrounding rock for different stress wave peak and stress wave frequency are compared and analyzed using UDEC software. Then, the combined support scheme of “secondary strong bolting and grouting” is proposed to control the severe deformation of the roadway. The primary support mainly adopts short hollow grouting anchor cable to strengthen the shallow broken surrounding rock within 0–3 m. The secondary reinforced support adopts long hollow grouting anchor cable or constant resistance large deformation anchor cable to fully mobilize the bearing capacity of deep surrounding rock and to ensure the long-term stability of the roadway. On-site monitoring showed that the scheme could effectively control the severe deformation of the roadway, which also provided a reference for other similar roadway support.

AbstractCoal bursts caused by mining-induced seismicity occur in the working face in steeply inclined and extra thick coal seams (SIETCSs). This article presents a case study in SIETCS through discrete element simulation as well as field investigation. The evolution of the coal burst was studied from the perspective of stress distribution, crack propagation and coal fragment ejection. The results showed that dynamic failure mainly occurred in the coal near the roof. The deep bottom coal and the top coal near the roof are under high static stress due to the clamping effect of the roof and floor and the rotation of the deep roof. The coal burst in the working face was triggered by mining-induced seismicity. The coal burst occurred when the sum of the static stress and the dynamic stress caused by the mining-induced seismicity exceeds the sum of the ultimate strength of the coal and the support strength, which was called the ‘clamping-rotation-dynamic loads’ mechanism. During the coal burst, the cracks developed rapidly into the deep bottom coal and the top coal on the roof side. The results can serve as an important reference for the coal burst failure mechanism in the working face in SIETCSs.

Copyright of Journal of Engineering Geology / Gongcheng Dizhi Xuebao is the property of Journal of Engineering Geology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Fractures generated in the overlying rock of coalfield fire areas can develop into air leakage and oxygen supply channels, thus helping sustain and intensify the fire. However, the formation, evolution, and distribution of such fractures are complex. This study used physical similarity simulation and numerical simulation to analyse the formation and evolution process of the underground coal field fire area at the Tanyaoqu coalmine. The first air leakage and oxygen supply channel in a coal fire area is formed when the overlying rock fractures reach the surface. The average collapse step distance of overlying rock after the formation of the first air leakage and oxygen supply channel in the research area was approximately two times that before. The length of the main fracture in the overlying rock increased gradually with coal combustion temperature. The angle between the main fracture and the horizontal plane decreased as temperature increased before the formation of the first channel and increased with temperature after the formation. When the coal combustion temperature was 600, 800, and 1000 °C, the distances from the coal seam damage position to the first air leakage and oxygen supply channel position are 122.7, 114.7, and 103.1 m, respectively. [ABSTRACT FROM AUTHOR]

Incorrect estimation of undercut dimensions in the block caving method can lead to the cessation of caving operations and loss of a large portion of deposits. Numerical modeling is one of the methods for determining the minimum caving span. Numerical and physical modeling methods are useful for an accurate understanding of caving operations. Accordingly, this research focused on investigating the performance of physical and numerical modeling in determining the effects of depth and joint orientation on the minimum required caving span for the initiation and propagation of caving. The physical model was made with 1.5*1.5 square meter dimensions and consisted of travertine blocks with 4*4 square centimeter dimensions. In addition, joints were modeled with dips of 0, 90, 45, 135, 30, and 120 degrees. The physical model could simulate ground stress conditions to great depths and show the behavior of the jointed rock mass in a two-dimensional space. Further, by capturing this behavior, it was possible to compare its result with UDEC software. The results demonstrated that the number of falling blocks and the height of the caving increased by increasing the dip. Furthermore, the formation of arches due to high horizontal stress stops the caving, which will occur again with the increasing span. Although the horizontal stresses and geometrical properties of the joints affect the shape of the caving area, its shape largely follows the dip and orientation of the rock mass joints. Poor draw control causes caved ore columns, which can lead to the formation of a stable arc. Finally, the height of the caved back increases in each span by increasing the depth while decreasing the dip of the joints.

Highlights: (1) Energy-absorption capacity of yielding rockbolts in rockbursts is evaluated. (2) Rock deformation and rockburst damage are quantitatively assessed. (3) Performance of yielding rockbolts in weak and strong rockbursts is studied. (4) Effects of floor support and cable bolts on controlling rockbursts are discussed.

The collapse of overlying rocks caused by the instability of residual coal pillars during lower coal seam mining significantly impacts its safety. This paper focuses on the gentle dipping coal seam group and utilizes the discrete element method (DEM) as the basis to comprehensively consider multiple factors through orthogonal experiments. In so doing it reveals the influence mechanisms of various factors on mining at a close distance under a residual coal pillar. Firstly, the process of lower coal seam mining under residual coal pillars in gently-dipping coal seams was simulated and analyzed based on a case study at the Baoping coal mine. Comparing the evolution characteristics of coal–rock fractures, stress changes, and displacement changes during the mining process reveals the mechanism of the joint instability of the lower coal seam, interlayer rock, coal pillars, and overlying strata under the disturbed conditions of lower panel mining. Secondly, an orthogonal simulation experiment was established using the width of the coal pillar and the thickness of the lower coal seam as variables. By comparing the development process of cracks, stress distribution, and rock displacement under different conditions, the research results indicate that the width of coal pillars has an impact on the maximum amount of coal pillar subsidence, while the thickness of the underlying coal seam has an impact on the time of subsidence, when hd ≥ 4.2 m (hj/hd ≤ 4) and w ≤ 14 m (w/hm ≤ 2), a large-scale collapse of the overlying strata of the coal pillar occurs. [ABSTRACT FROM AUTHOR]

Aiming at the problem that it is difficult to popularize in Guizhou mining area due to high filling cost, taking 11071 working face of Panzhihua Coal Mine in Liupanshui as the research background, the idea and technology of strip interval filling mining are put forward. Through theoretical analysis and mechanical tests, the reasonable range of strip spacing distance and mechanical parameters of filling body were obtained. The FLAC3D simulation software is used to analyze the stress field, displacement field, and plastic zone of five kinds of design schemes, and the UDEC simulation software is used to carry out the filling mining of No.7 coal seam. The results show that the strip interval filling mining technology regards the empty roof area as a controllable underground spatial structure, and the single spatial stress field changes little. The load on the immediate roof is gradually transferred from the top of the empty roof area to the top of the filling body. The “filling body-direct roof” structure improves the self-bearing capacity of the immediate roof, and the overlying surrounding rock migration is controlled. With the increase of mining depth, it gradually tends to the original rock stress, and the control effect on surface subsidence is more significant. Finally, “filling 3 m interval 3 m” is determined as the optimal filling scheme. In the process of simulated filling mining, the peak stress in the stress concentration area of the front coal wall shows a trend of “increase-decrease-increase,” and the peak stress curve of the immediate roof in the middle of the stope changes from “increase-decrease ‘trend to’ increase-decrease-increase-decrease” trend. The rock layer near the immediate roof is in a stress concentration state in the coal wall area on both sides, and the middle part is not obvious. [ABSTRACT FROM AUTHOR]

Uniaxial compression tests on Hydro-StoneTB® porous cubes were numerically modeled using a distinct element program (UDEC) for two-dimensional geological material simulation. Under uniaxial compression, numerical square models with varying void porosity and geometry, identical to an experimental testing, were simulated. The open-ended longitudinal openings in the experimental specimens, cubic porous Hydro-StoneTB®, were of various forms, sizes, uniformity, and distributions. UDEC's Voronoi tessellation joint generator was used to simulate the material of the experimental specimens as an assembly of discrete units interacting along their boundaries. The results showed that the numerical results from the two-dimensional numerical analysis utilizing the discrete element method, UDEC program, were consistent with the experimental data. However, when compared to the experimental values, the numerical analysis produced conservative values for both uniaxial compressive strength and Young's modulus due to the modeling of a three-dimensional material in a two-dimensional planar strain. Furthermore, regardless of void porosity or geometry, the tension (axial splitting) failure mode is the dominating failure mode of rock-like materials including voids, similar to the experimental results. It was also demonstrated that using a power relationship, two-dimensional numerical data can be converted to three-dimensional experimental results. [ABSTRACT FROM AUTHOR]

With the increase in mining depth, coal bursts have become a major challenge in the safe mining of steeply inclined and extra thick coal seams (SIETCSs). Based on a typical mining induced seismicity triggered coal burst (MSTCB) in SIETCS, a large-scale numerical model was developed using the Universal Distinct Element Code. The numerical model was calibrated and validated by laboratory results and field observations. The stress evolution, crack development and ejection velocity patterns in the MSTCB were analysed, and the effect of mining induced seismicity vibration velocity on the MSTCB was discussed. The results show that a triangular static stress concentration zone is formed in the coal on the roof side. And the high-energy mining induced seismicity leads to high dynamic stresses in the coal at the roof side rib and top of the headentry. Coal bursts occur under the superposition of static and dynamic stresses. The MSTCB results in tensile failure near the headentry surface and shear failure in the depth. The vibration velocity has a significant effect on the roof side rib and top of the headentry, while it has only a slight effect on the working face rib and bottom of the headentry. The dynamic stress and ejection velocity in the roof side rib and top of the headentry are positively correlated with the vibration velocity. Finally, measures for MSTCB prevention were proposed. The findings presented in this study can provide guidance for the prevention and control of MSTCBs in SIETCSs.

Underground mining activity in the karst mountain in southwestern China has induced several large-scale rocky landslides and has caused serious casualties. At present, there is a lack of systematic research on the formation mechanism of landslides in this area using multi-method fusion technology. First, the orthophoto images of the landslide area obtained by UAV photography were used to analyze the deformation characteristics of the landslide. Second, the failure characteristics of the strata overlying the goaf were analyzed by geophysical detection. Finally, the deformation response characteristics of the mountain under underground mining were analyzed by UDEC numerical simulation. The results revealed that during the underground mining, the failure process of the mountain occurred in four stages: fracture expansion, subsidence and collapse, shear sliding, and multi-level sliding. Gently dipping soft–hard alternant strata and a blocky rock mass structure formed the geological foundation of the landslides. Underground mining accelerated the fracturing of the overlying strata and the formation of a stepped penetrating sliding surface. Tensile movement of the structural planes of hard sandstone in the free face, and shear sliding of the weak mudstone layer, were the main causes of the landslides. The slope instability mode was tension-shear fracturing, shear sliding, back toppling, and compressive shear failure. In addition, the fracture propagation in the overlying strata and damaged geological structure revealed by the geophysical detection were consistent with the simulation results. This study provides ideas for the precise countermeasures of disaster prevention and mitigation for similar landslides in this area. [ABSTRACT FROM AUTHOR]

The assessment of yielding rockbolt performance during rockbursts with actual seismic loading is essential for rockburst supporting designs. In this paper, two types of yielding rockbolts (D-bolt and Roofex) and the fully resin-grouted rebar bolt are modeled via the "rockbolt" element in universal distinct element code (UDEC) after an exact calibration procedure. A two-dimensional (2D) model of a deep tunnel is built to fully evaluate the performance (e.g., capacity of energy-absorption and control of rock damage) of yielding and traditional rockbolts based on the simulated rockbursts. The influence of different rockburst magnitudes is also studied. The results suggest that the D-bolt can effectively control and mitigate rockburst damage during a weak rockburst because of its high strength and deformation capacity. The Roofex is too "soft" or "smooth" to limit the movement of ejected rocks and restrain the large deformation, although it has an excellent deformation capacity. The resin-grouted rebar bolt can maintain a high axial force level during rockbursts but is easy to break during dynamic shocks, which fails to control rapid rock bulking or ejection. Three types of rockbolts cannot control the large deformation and mitigate rockburst damage effectively during violent rockbursts. The rockburst damage severity can be significantly reduced by additional support with cable bolts. This study highlights the effectiveness of numerical modeling methods in assessing the complex performance of yielding rockbolts during rockbursts, which can provide some references to improve and optimize the design of rock supporting in burst-prone grounds. [ABSTRACT FROM AUTHOR]

Incorrect estimation of undercut dimensions in the block caving method can lead to the cessation of caving operations and loss of a large portion of deposits. Numerical modeling is one of the methods for determining the minimum caving span. Numerical and physical modeling methods are useful for an accurate understanding of caving operations. Accordingly, this research focused on investigating the performance of physical and numerical modeling in determining the effects of depth and joint orientation on the minimum required caving span for the initiation and propagation of caving. The physical model was made with 1.5*1.5 square meter dimensions and consisted of travertine blocks with 4*4 square centimeter dimensions. In addition, joints were modeled with dips of 0, 90, 45, 135, 30, and 120 degrees. The physical model could simulate ground stress conditions to great depths and show the behavior of the jointed rock mass in a two-dimensional space. Further, by capturing this behavior, it was possible to compare its result with UDEC software. The results demonstrated that the number of falling blocks and the height of the caving increased by increasing the dip. Furthermore, the formation of arches due to high horizontal stress stops the caving, which will occur again with the increasing span. Although the horizontal stresses and geometrical properties of the joints affect the shape of the caving area, its shape largely follows the dip and orientation of the rock mass joints. Poor draw control causes caved ore columns, which can lead to the formation of a stable arc. Finally, the height of the caved back increases in each span by increasing the depth while decreasing the dip of the joints. [ABSTRACT FROM AUTHOR]

The deterioration of rock mass in the hydro-fluctuation belt and reservoir-induced earthquakes in the Three Gorges Reservoir Area (TGRA) have great impact on the stability of bedding rock slope. The cumulative damage mechanism and stability of bedding rock slope under the deterioration of rock mass in the hydro-fluctuation belt, when subjected to repeated seismic loads, were investigated using the shaking table model test and Universal Distinct Element Code (UDEC) numerical simulation. Under the continuous action of seismic loads, the peak ground acceleration (PGA) amplification coefficient clearly weakens; the cumulative displacement, pore water pressure and earth pressure of the slope show a change trend of increase, increase and decrease, respectively; the damping ratio and damage degree (DG) of the slope increase, while the natural frequency of the slope decreases gradually; the nonlinear cumulative damage mechanical models of the slope in the stages of microseism—small earthquake action and strong earthquake action can be characterized by the cubic function of “S-type” and the exponential function of “steep rise type,” respectively. Based on the good agreement between the experimental observations and numerical simulation results, the evolution process of the dynamic cumulative damage—instability and typical failure modes of the slope (including the hydro-fluctuation belt) was observed. Furthermore, the stability of the slope increases with the decrease of the slope height (A), slope angle (B), seismic load amplitude (C), seismic load frequency (D), or deterioration depth of rock mass in hydro-fluctuation belt (E), while it decreases with the increases of the strength degradation ratio of the rock discontinuities in hydro-fluctuation belt (F); the ranking of sensitivity of the mentioned-above influence factors is C > A > D > E > B > F according to the orthogonal analysis method (OAM). [ABSTRACT FROM AUTHOR]

Block identification serves as a critical preprocessing step in rock block system mechanical analysis. Presently, research primarily focuses on refining identification algorithms to accommodate increasingly complex joint models or enhance computational efficiency, while overlooking the influence of the block system expression derived from identification on subsequent mechanical analysis. Specifically, the precise role of small blocks within a block system during numerical simulation remains uncertain. Consequently, this paper developed a novel two-dimensional block identification method integrating small block treatment. The essence of this treatment lies in eliminating the short edges shared between small blocks and their adjacent blocks. Here we find that small block treatment is crucial for enhancing the efficiency and convergence of mechanical simulation. Through comparisons with the existing treatment methods using three cases in the Universal Distinct Element Code (UDEC), our method demonstrates superior computational efficiency and exhibits the best consistency with the results of the original models in Cases 1 and 2. Moreover, our method is uniquely capable of successfully conducting numerical simulations for Case 3. We anticipate our study to be a starting point for advancing the geometric optimization of block systems and bridging geometric and mechanical analyses in the realm of rock block mechanics analysis. [ABSTRACT FROM AUTHOR]

In this work, the thermo-mechanical stress–strain history of an Alpine slope is analyzed, with particular focus on the historical Cimaganda large landslide (Sondrio Province, Italy), which mobilized an estimated volume of 7.5 mm3 of rock material. Accurate geomorphological and geomechanical characterization involving field surveys and laboratory testing was carried out, leading to the development of a conceptual model of the slope. A thermo-mechanical semi-coupled approach was developed, considering both glacial debuttressing and thermo-mechanical effects due to gradual exposure of the slope to atmospheric conditions and paleo-temperature redistribution resulting from the Last Glacial Maximum deglaciation. A 2D distinct-element numerical approach was adopted, supported by a 2D finite-element analysis to simulate heat diffusion over the Valley cross-section. Modelling results allow to simulate the general evolution of the Cimaganda rock-slope and to highlight the significance of thermal processes in preparing rock-slope instabilities. While the mechanical effect of ice thickness reduction alone brings about moderate rock mass damage, the introduction of temperature couplings results in a substantial increase of damage, representing a significant factor controlling the stress–strain evolution of the slope. Simulated displacement and the development of a deep region of shear strain localization at a depth roughly corresponding to that of the detected Cimaganda sliding surface, allow to highlight the significance of temperature influence in preparing the rock-slope to instability. Highlights : The effects of surface temperature variations on the stress-strain evolution of an Alpine rock-slope are examined through numerical simulation. This study contributes to the understanding of Thermo-Mechanical processes driving damage in rock masses. Results showed how Thermo-Mechanical stresses induced by temperature variations at both long- and short-term timescales, can represent a significant preparatory factor to rock slope instability. The factors introduced in the analysis allowed to simulate the general evolution of the Cimaganda rock-slope, which is corroborated by field survey data. [ABSTRACT FROM AUTHOR]

احداث پرده آب‌بند یکی از روش‌های متداول کنترل تراوش از زیر پی سدها است. به منظور ایجاد پرده آب‌بند، تعدادی چال حفر شده و در این چال‌ها عملیات تزریق صورت می‌گیرد. در عملیات تزریق درزه‌های موجود در توده‌سنگ با استفاده از دوغاب مناسب پر می‌شود. از آنجا که نفوذپذیری ماده‏سنگ بسیار کم و قابل صرفنظرکردن است، عمق نفوذ دوغاب به پارامترهای متعددی از جمله خصوصیات دوغاب، فشار تزریق و مشخصات ناپیوستگی‌ها بستگی دارد. استفاده از روش‌های عددی برای تعیین آرایش بهینه گمانه‌های تزریق باعث صرفه‌جویی در هزینه‌ها و کاهش زمان تزریق می‌شود. در این مقاله با توجه به خصوصیات توده‌سنگ در محدوده احداث سد سردشت با استفاده از نرم‏افزار المان مجزای UDEC آرایش بهینه سیستم آب‌بندی پی سد بررسی شده‌است. براساس نتایج این تحقیق عمق بهینه پرده 40 متر، فاصله‌داری چال‌های تزریق در پی و تکیه‌گاه‌ها به ترتیب 3 متر و 5 متر، پیشنهاد می‌شود. همچنین زاویه انحراف بهینه چال‌ها 17 درجه برآورد شده‌است. مقایسه نتایج بدست آمده از مدل‌سازی عددی با مقادیر اندازه‏گیری شده نشان‌دهنده قابلیت بالای روش عددی در برآورد فشار تزریق و تحلیل نشت از پی سد سردشت است.

Block-flexure is the commonest mode of toppling failure and can be frequently encountered in anti-inclined rock slopes. In this work, the failure mechanism of block–flexure toppling (BFT) was investigated using centrifugal and numerical models. The numerical model was configured using the Universal Distinct Element Code (UDEC) and calibrated with the results of the centrifuge test. All the simulation results, including the measured displacements, failure load, and failure surface, are generally in line with that of experimental results. Further, the results of simulations show that, well before the appearance of instability in the jointed rock slope, slipping failure and opening fractures had occurred in the interlayer. Moreover, all the acting points of normal forces in steep joints are located between the bottoms and the midpoints of the columns under consideration, in the process of toppling failure. Finally, sensitivity of joint parameters, including the connectivity rate of discontinuous cross-joints, the thickness of the rock column joint, joint friction angle, and joint cohesion, were performed to investigate the effects of the tensile strength of intact rock. The results indicate that joint cohesion and thickness of the rock column greatly influence the failure load (represents safety factor of the slope). The connectivity rates of the discontinuous cross-joints and joint friction angle were found to have significant effects on the shape and location of the basal failure plane. This research would provide a deep understanding on the failure mechanism of BFT for relevant scholars. [ABSTRACT FROM AUTHOR]

UDEC is used to analyze the surface subsidence characteristics under different steps of fully mechanized and layered mining face. The research results show that the trend of surface subsidence in the lower layer mining face and the fully mechanized mining face is basically the same, and the extreme value is in the middle back position of the goaf, but the maximum subsidence value of the lower layer face increases by about 30.9%. The sinking coefficient is reduced by about 11.5%; the extreme value of sinking speed of fully mechanized mining face increases with the increase of step of working face, and shifts to the direction of push mining, while the extreme value of sinking speed of lower layered working face relatively decreases; the horizontal movement of the stratum mining face is more violent and fluctuating than that of the fully mechanized mining face. The positive and negative values of the horizontal movement increase by approximately 45.2% and 63.6%, respectively. The influence range of the fully mechanized mining face is greater than that of the layered mining face. Compared with the fully mechanized mining face, the leading influence angle of layered mining under the layer mining face is reduced by about 4.38%, and the boundary angle is reduced by about 1.43%. The conclusion is to provide guidance for rational determination of the mining method.

The shear characteristics of rock joints under constant normal stiffness (CNS) boundary condition are critical to the stability of underground engineering rock mass. Five different rock joint profiles with random morphology were constructed using the independent segmentation method of the Hurst exponent. Based on the continuously yielding joint model, the discrete element calculation models of double rough parallel joints with different joint surface roughness and spacing of joints under CNS boundary condition were established using UDEC software to investigate the shear effect of joints under CNS boundary condition. The results show that under CNS boundary condition, the shear stress, normal displacement, normal stress, and surface resistance index (SRI) all increase with the increase of joint surface roughness coefficient (JRC) for both single- and double-joint specimens, and these of the single-joint are greater than these of the double-joint. With the increase of the joint spacing d, the peak shear stress τy, and the peak surface resistance index (SRIp) show a gradually increasing trend, and the influence of d on τy and SRIp is greater with the increase of JRC, and both τy and SRIp show a linear increasing trend with the increase of JRC. With the increase of d for the same JRC, both normal displacement and normal stress show a gradually increasing trend, and also, the increased amplitudes gradually increase; with the increase of JRC for the same d, the two parameters also show an increasing trend.

Based on the two-dimensional discrete element software UDEC, this article studied the dynamic response laws of rock slopes with inverse nonpersistent joints by the combination of different dip angles and spacing of joint and length of rock bridge under earthquake. The results showed that the existence of the joint surface had a significant impact on the dynamic response of the slope. When the dip angle of inverse nonpersistent joint increases or the joint spacing decreases, the PGA amplification factor of each monitoring point on the slope surface increases, and the influence range is mainly concentrated in the middle of the slope surface to the slope shoulder. Along the horizontal direction of the slope, closer to the shoulder of the slope, the PGA amplification factor increases with the increase of the joint dip angle and the decrease of rock bridge length and joint spacing; along the vertical direction, the PGA amplification coefficient curve increases first, then decreases, and then increases with the change of elevation, and the dynamic response of slope reacts the most violent where the joints are densely developed.

Underground mining activity in the karst mountain in southwestern China has induced several large-scale rocky landslides and has caused serious casualties. At present, there is a lack of systematic research on the formation mechanism of landslides in this area using multi-method fusion technology. First, the orthophoto images of the landslide area obtained by UAV photography were used to analyze the deformation characteristics of the landslide. Second, the failure characteristics of the strata overlying the goaf were analyzed by geophysical detection. Finally, the deformation response characteristics of the mountain under underground mining were analyzed by UDEC numerical simulation. The results revealed that during the underground mining, the failure process of the mountain occurred in four stages: fracture expansion, subsidence and collapse, shear sliding, and multi-level sliding. Gently dipping soft–hard alternant strata and a blocky rock mass structure formed the geological foundation of the landslides. Underground mining accelerated the fracturing of the overlying strata and the formation of a stepped penetrating sliding surface. Tensile movement of the structural planes of hard sandstone in the free face, and shear sliding of the weak mudstone layer, were the main causes of the landslides. The slope instability mode was tension-shear fracturing, shear sliding, back toppling, and compressive shear failure. In addition, the fracture propagation in the overlying strata and damaged geological structure revealed by the geophysical detection were consistent with the simulation results. This study provides ideas for the precise countermeasures of disaster prevention and mitigation for similar landslides in this area.

There have been reported cases of wellbore instability near-wellbore area due to failures associated with drilling operations, the magnitude and extension of the failure area is imprecisely evaluated by copious researchers due to some of the limitations posed by the numerical simulations used. Therefore, this paper estimates a failure zone area near-wellbore based on the conducted coupled Thermal-Hydro Mechanical (THM) analysis integrated into UDEC, wellbore enlargement, and Failure area analysis. A two-dimensional horizontal plane stress model of 2 × 2 m is created using rock properties from a vertical well, and a wellbore radius of 0.11 m is drilled under confined stress using the in-situ stresses. The Voronoi tessellation joint system was used to create randomly polygon fractures referred to as Discrete Fracture Size (DFS). The coding of various parameters was performed to study the near-wellbore stability for both underbalanced and overbalanced drilling conditions, the four cases of DFS were used against seven wellbore pressures. After comparing with other DFS, The parametric analysis at DFS of 0.1 and 0.12 m offers low shear and block displacements that led to an induced shear and tensile failures, respectively. However, to evaluate the optimal DFS, a comparison is done using Failure area and Borehole enlargement analysis, and the results show that DFS of 0.12 m yields significantly low wellbore enlargement rate and failure area zone for all stress ratios and drilling cases. The proposed simple model for wellbore failure area estimations is verified against a well vertical section located in Northern China.

There have been reported cases of wellbore instability near-wellbore area due to failures associated with drilling operations, the magnitude and extension of the failure area is imprecisely evaluated by copious researchers due to some of the limitations posed by the numerical simulations used. Therefore, this paper estimates a failure zone area near-wellbore based on the conducted coupled Thermal-Hydro Mechanical (THM) analysis integrated into UDEC, wellbore enlargement, and Failure area analysis. A two-dimensional horizontal plane stress model of 2 × 2 m is created using rock properties from a vertical well, and a wellbore radius of 0.11 m is drilled under confined stress using the in-situ stresses. The Voronoi tessellation joint system was used to create randomly polygon fractures referred to as Discrete Fracture Size (DFS). The coding of various parameters was performed to study the near-wellbore stability for both underbalanced and overbalanced drilling conditions, the four cases of DFS were used against seven wellbore pressures. After comparing with other DFS, The parametric analysis at DFS of 0.1 and 0.12 m offers low shear and block displacements that led to an induced shear and tensile failures, respectively. However, to evaluate the optimal DFS, a comparison is done using Failure area and Borehole enlargement analysis, and the results show that DFS of 0.12 m yields significantly low wellbore enlargement rate and failure area zone for all stress ratios and drilling cases. The proposed simple model for wellbore failure area estimations is verified against a well vertical section located in Northern China. [ABSTRACT FROM AUTHOR]

High pre-tension bolt is an effective strata control technique and is the key to ensure the stability of anchorage and roadway. Based on the performances of high energy storage tension rock bolts in different rock properties, this study proposed a constitutive model to describe the energy balance of anchor under uniaxial compression. UDEC was used to simulate the behaviour of anchor in coal under uniaxial compression and the results were analysed to study the rock mechanical properties, degree of damage and energy evolution. Simulation results showed that tension rock bolts can improve the mechanical properties and energy storage capacities of the anchor. The energy evolution was divided into three stages: (i) the external work was stored in the form of elastic strain energy (Ue) in the anchor prior to the yielding strength; (ii) the elastic strain energy reached its maximum near the peak strength; (iii) energy was dissipated from fracture friction (Wf), plastic deformation (Wp) and acoustic emission (Ur) during post-peak stage. The installation of tension rock bolts was more suitable for medium hard rock (e.g. sandy mudstone), whereas it was not effective for hard rock (e.g. sandstone).

As a result of the mining of a C3 coal seam in a mine in Guizhou, perilous rock masses on the surface collapsed. In this study, the stability of perilous rock masses on the surface of the coal mine before and after mining was calculated and examined, and the movement law of the overlying strata in the goaf, the movement and deformation law, and the failure mode of perilous rock were analyzed. This study provides a theoretical basis for the treatment of unstable rock and coal seam mining, and has important guiding significance for the safe and efficient production of the mine. The results show that: (1) The perilous rock is in a basically stable state without the influence of mining. Through theoretical analysis and the construction of the collapse model of perilous rock, it is judged that perilous rocks W1, W3, W4, and W7 were basically stable, perilous rocks W2 and W5 were in an unstable state, and perilous rock W6 was stable without heavy rainfall. (2) As a result of the mining of the C3 coal seam, the cracks in the upper strata began to develop to the surface, and the longitudinal separation cracks gradually appeared between the surface perilous rock and the rock matrix. Due to the existence of these cracks, the perilous rock had a downward shear force. In addition, due to the heavy rainfall in the Guizhou area, the transient saturated zone of perilous rock is expanding and the strength of perilous rock is reduced. The seepage increases the sliding force of the perilous rock and aggravates the opening of cracks at any time. (3) The stability of the surface perilous rock mass is largely affected by the mining of underground coal mines. The simulation analysis was repeated using the method of setting coal pillars. When 45 m permanent protection coal pillars are set at both ends, and 15 m local protection coal pillars are set at 60 m, the safety of coal mining can be ensured without affecting the surface and perilous rock.

This paper aims to study the mechanical behaviour and failure mechanism of intact rocks under different loading conditions using the grain based model implemented in the universal distinct element code (UDEC). The grain based numerical model is a powerful tool to investigate complicated micro-structural mechanical behaviour of rocks. In the UDEC grain based model, the intact material is simulated as assemblies of a number of polygonal blocks bonded together at their contact areas. To investigate the ability of such a numerical framework, uniaxial and triaxial compression tests as well as direct tensile test were simulated in UDEC and then the results were compared with the laboratory experiments undertaken on Hawkesbury sandstone. There was a good agreement between the experimental and numerical under different loading conditions. In order to investigate the effect of micro-properties of the grain based model, blocks and contacts, on the laboratory scale intact rocks, a set of parametric study was undertaken. The results from this analysis confirmed that the block size is an intrinsic characteristic of a model which has significant effects on the mechanical behaviour of the numerical models. Also, it was concluded that the cohesion and friction angle of contact surfaces control both uniaxial and triaxial compressive strengths. Finally, it was found that in the triaxial compression test, as the applied confining pressure increases, the effect of contact cohesion on the strength decreases while the effect of friction angle increases.

طراحی سازه‌های زیرزمینی باید با توجه به شرایط ساختاری توده‌سنگ صورت گیرد. در این زمینه شرایط قرارگیری درزه‌ها، زوایه آنها نسبت به فضای زیرزمینی و زاویه درز‌ها نسبت به هم از اهمیت ویژه‌ای برخوردار است. در این مطالعه پایداری ساختاری تونلی با مقطع نعل اسبی در توده‌سنگ درزه‌دار حاوی دو دسته درزه با فاصله‌داری‌های متفاوت مورد بررسی قرار گرفت. برای مدلسازی درزه‌ها پارامترهای سختی برشی درزه‌ها، شیب دسته درزه‌ها، بازشدگی و فاصله‌داری، نسبت فاصله‌داری درزه‌ها به عرض دهانه تونل () و نسبت تنش‌های برجا به عنوان متغیرهای مدلسازی در نظر گرفته شد. مساحت زون ریزش با تغییر هر یک از شاخص‌ها به طور جداگانه محاسبه شد و بهترین حالت ممکن که دارای کمترین سطح زون ریزش است، تعیین شد. نتایج نشان داد که اگر اختلاف شیب دو دسته‌درزه کمتر از 40 درجه باشد بیشترین سطح زون ریزش تشکیل می‌شود و افزایش اختلاف شیب دسته درزه‌ها بیش از 40 درجه موجب کاهش سطح ریزش می‌شود. همچنین افزایش نسبت فاصله‌داری درزه‌ها به عرض دهانه تونل منجر به افزایش مساحت زون ریزشی و ناپایداری بیشتر می‌شود. ناپایداری در نسبت مذکور در بازه‌های 11/0 تا 33/0 وجود دارد و برای مقادیر بیش از 33/0 برای نسبت ذکر شده به علت افزایش ابعاد بلوک و عدم امکان ریزش، ناپایداری به صفر می‌رسد. ایمن‌ترین حالت و کم‌ترین مساحت زون ریزش در نسبت و بیش‌ترین مساحت در نسبت اتفاق افتاد.

Uncertainty in input fracture geometric parameters during analysis of the stability of jointed rock slopes is inevitable and therefore the stochastic discrete fracture network (DFN) — distinct element method (DEM) is an efficient modeling tool. In this research, potentially unstable conditions are detected in the right abutment of the Karun 4 dam and downstream of the dam body as a case study. Two extreme states with small and relatively large block sizes are selected and a series of numerical DEM models are generated using a number of validated DFN models. Stability of the rock slope is assessed in both static and dynamic loading states. Based on the design basis earthquake (DBE) and maximum credible earthquake (MCE) expected in the dam site, histories of seismic waves are applied to analyze the stability of the slope in dynamic earthquake conditions. The results indicate that a MCE is likely to trigger sliding of rock blocks on the rock slope major joint. Furthermore, the dynamic analysis also shows a local block failure by the DBE, which can consequently lead to slope instability over the long term. According to the seismic behavior of the two models, larger blocks are prone to greater instability and are less safe against earthquakes. [ABSTRACT FROM AUTHOR]

A series of DEM-FDM coupled numerical simulations were performed to study the failure process of the isolated stone column in soft soil according to the Universal Distinct Element Code. In the proposed numerical model, the stone column was represented by a rigid blocky system based on the Voronoi tessellation; the surrounding soil was represented as the continuum Mohr-Coulomb material. The settlement, bulging deformation, and failure processes were tracked during the complete loading process, which enables the column-soil interactions to be observed visually. The deformation process and corresponding failure modes of isolated stone columns embodied in soft soil are graphically illustrated according to the numerical analysis. The traditional ultimate bearing capacity prediction formulas for the isolated stone column cannot be directly used for the local shear failure mode. The soil stiffness should be the critical parameter to calculate the lateral confining stress rather than the soil shear strength. [ABSTRACT FROM AUTHOR]

Laboratory tests showed that the average direct tension strength (DTS0) of the Lac du Bonnet granite was 6.9 MPa, and significantly less than its average Brazilian tension strength (BTS0), i.e., 8.8 MPa. In this paper, a grain-based model (GBM) with a 2D universal discrete element code (UDEC) was used to explore factors that could control the difference in the direct tension strength (DTS) and Brazilian tension strength (BTS) of the Lac du Bonnet granite from the micromechanical point of view. Micro parameters of the grain contacts were investigated, i.e., the tensile strength (T), cohesion (c), friction angle (φ), and the ratio of the shear to normal stiffness (λ). Results show that both BTS and DTS highly depended on log(λ). With an increase in log(λ), the BTS first increased nonlinearly and decreased instead after achieving a peak while the DTS always increased nonlinearly. The BTS was sensitive to T and c, while the DTS was mainly sensitive to T. The ratio of DTS to BTS first decreased with the increase of log(λ) and then increased instead from a minimum. It was found that the GBM with UDEC can well reproduce the uniaxial compression, direct tension and Brazilian tests of the Lac du Bonnet granite simultaneously only on condition that λ = 2.51. This indicated that the shear stiffness of the grain boundary in the Lac du Bonnet granite may be several times larger than its normal stiffness. This may be the main factor controlling the difference in the BTS and DTS of the Lac du Bonnet granite. [ABSTRACT FROM AUTHOR]

تحلیل پایداری شیب دیواره­های معادن سطحی به­خصوص شیب­های سنگی، از جمله مهم­ترین مسائل مورد توجه در طراحی معادن سطحی است که با اجرای یک مدل­سازی ژئوتکنیکی قابل دسترسی است. این مدل ژئوتکنیکی برای توده­سنگ­های درزه­دار ایجاد می­شود و شامل دو زیر مدل به نام­های مدل هندسی و مکانیکی است. مدل هندسی توده­های سنگی شامل شبکه­ی شکستگی­های مجزا (DFN) است که براساس وضعیت توپوگرافی، حدود مدل، بلوک­های تکتونیکی و ناپیوستگی­های موجود ایجاد می­شود. مدل مکانیکی با هدف تحلیل­های مکانیکی (استاتیکی یا دینامیکی) توده­سنگ ایجاد شده و شامل بررسی پارامترهای ژئومکانیکی، وضعیت تنش‌های برجا، شرایط آب­های زیرزمینی، قوانین حاکم بر مقاومت و رفتار توده­سنگ، روش­های تحلیل پایداری و در نهایت قیاس و نتیجه­گیری در مورد پایداری مکانیکی است. هدف از این مقاله، ایجاد یک مدل ژئوتکنیکی و تحلیل مکانیکی مربوط به آن برای بلوک تکتونیکی شماره­ی 2-4 معدن سنگ آهن چغارت در دو مقطع از کاواک قدیم و جدید و مقایسه­ی آن­ها برای بررسی تغییرات پایداری دیواره­ی مذکور در طراحی شیب جدید است.برای دستیابی به این هدف، مدل­سازی هندسی با روش شبکه­ی شکستگی­های مجزا در سیستم ناپیوستگی­های نامحدود غیر ترتیبی با تمرکز بر حذف اثر تصادفی مدل­سازی هندسی، و تحلیل مکانیکی مربوط به آن به­وسیله­ی نرم­افزار المان مجزای UDECانجام و نتایج تحلیل شدند. بررسی­ها نشان دهنده­ی جابجایی­های کمتر و ثبات بیشتر دیواره­ی جدید نسبت به دیواره­ی قدیمی است.

Stress arching effect during the excavation of broken surrounding rock in underground engineering has an important influence on the stability of surrounding rock after underground excavation. To determine the stress arching effect in horizontally layered jointed rock mass, the stress arching characteristics of surrounding rock mass after excavation is analyzed in this study by using a series of numerical tests. The formation mechanism of stress arch is revealed through a comparison of the stress characteristics of a voussoir beam structure and theoretical analysis of multi-block mechanical relationship of jointed rock mass. The method for determining the boundaries of a stress arching zone is proposed, and the influence of various factors on a stress arch is further discussed. Results show that after the excavation of horizontally layered jointed rock mass, the stress arch bunch (SAB) is formed in the lower strata above the cavern, and the global stress arch (GSA) is formed in the higher strata, both of which are symmetrical arch stress patterns. The SAB is the mechanical manifestation of the voussoir beam structure formed by several low-level sandstone layers, and the GSA is caused by the uneven displacement between blocks. Compared with the GSA, the SAB is more sensitive to various influencing factors. The extent of stress arching zone decreases with the increase of an internal friction angle of the joint, lateral pressure coefficient, and overburden depth. In addition, the joint spacing of rock strata is conducive to the development of a stress arch. Results can provide technical support for deformation control and the stability analysis of broken surrounding rock in underground engineering.