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Simulation of roof shear failure in coal mine roadways using an innovative UDEC Trigon approach
Abstract
Shear failure is a common failure mechanism in underground coal mine roadways. This paper presents an innovative numerical approach to simulate shear failure of a coal mine roadway roof. The distinct element code, UDEC, incorporating a proposed Trigon logic is employed for the study. Using this approach, shear failure in the mine roof characterized by fractured initiation and propagation is successfully captured. The results suggest that shear failure of the roadway roof initiates at the roadway corners and then progressively propagates deeper into the roof, finally forming a large scale roof failure. The numerical results confirmed the time sequence of marked microseismic activity, significant stress changes and accelerated displacement during the process of a roof fall. The effect of rock bolting in the control of roof shear failure in a roadway is evaluated using the UDEC Trigon approach. It is found that the installation of rock bolts constrains rock dilation, reduces failure of rock bridges and maintains rock strength thereby leading to a significant decrease in roof sag.
... Numerical simulation has been widely accepted for its unique repeatability and data diversity, especially Universal Distinct Element Code (UDEC). The newly added triangular element can effectively simulate the fracture expansion of coal/rock mass, and the embedded fish program can also complete the tracking and positioning of fractures [13,14]. Bai et al. [15] used a Discrete Element Method(DEM) investigation of the fracture mechanism of rock disc containing hole(s) and its influence on tensile strength and explored the effects of pore location, size, and quantity on its tensile strength and fracture. ...
... The bulk modulus (K) and shear modulus (G) of the blocks in the numerical model can be calculated with the following equation (13) and equation (14) [27]. Table 3 shows the microparameters of coal and rock parting blocks after calibration. ...
... Therefore, in order to quantitatively analyze the internal damage of CRCS, a fish function is used to record the total length of cracks as well as the length of shear and tensile cracks caused by stick-slip and instability of CRCS. The damage parameter (D) is proposed according to the analysis of Gao [13] and Wu [31]. ...
Unloading excavation can increase the possibility of rock burst, especially for coal seam with rock parting. In order to explore the evolution process of rock burst under lateral unloading, the combination of in situ measures and numerical experiments is used to study. The following four points were addressed: (1) the coal seam with rock parting easily causes the stick-slip and instability along the interface, and the process of stick-slip and instability has hysteresis characteristics; (2) the greater the degree of unloading or the smaller the interface friction angle of the Coal-Rock Parting-Coal Structure (CRCS), the more likely it is for stick-slip and instability to occur; (3) the abnormal increase of shear stress and slip dissipation energy can be used as the precursory information of the stick-slip and instability of CRCS; (4) the damage intensity of rock burst induced by stick-slip and instability of CRCS can be reduced by reducing the unloading speed or increasing the roughness of interface. The research results can be used for early warning and controlling of dynamic disaster induced by stick-slip instability in coal seam with rock parking.
... Under the infuence of the engineering excavation disturbance, the environment of the roadway tends to be complicated, and the risk potential of the roof falling and the rib spalling is increased [1]. On the one hand, roof falling and rib spalling seriously threaten the safe production of the mine; on the other hand, they block the roadway, restrict the operation of the production system, and seriously threaten the efcient production of the mine [2][3][4]. At the same time, the frequent roof falling forms ultrahigh roadway, and the support cost is huge, which seriously afects the beneft of the mine. ...
... We take the microsegment dξ at the distance |ξ| from the coordinate origin, and the load within the microsegment can be simplifed as the concentrated force, that is, dF � qdξ. Te stress generated by dF at any point M(x, z) of the lower rock layer can be solved according to formula (3). Here, the vertical and horizontal distances between the action point of dF and the point M are, respectively, expressed as |z| and |x − ξ|. ...
... 3) ...
Driven by the high original rock stress and engineering disturbance, the ultrahigh roadways of about 7∼8 m appear in the deep coal mine. Based on the field measured data, the stability in different zones of roof and rib is divided. From shallow to deep, the roof forms potential leakage and falling zone, metastable zone, and stable zone, and the coal rib forms potential spalling zone, crack developed zone, and transition zone. Based on the theory of transverse isotropy, the mechanical analysis model of layered roof is established, the real stress state at any point of layered roof is deduced, the roof failure criterion is constructed, and the theoretical solution of the failure depth of layered roof is obtained. The mechanical model of pillar bar of the splitting coal body in the potential spalling zone is established, and the deflection deformation expression of the pillar bar is derived. In view of the deformation and failure characteristics of the roof and rib, this paper puts forward the cooperative integrated support technology based on high-strength prestressed cables and deep-shallow hole grouting, which effectively reduce the hidden danger of roof falling and eliminate the formation of ultrahigh roadway from the source.
... He et al. (2015) pointed that if the walls of roadway have no supports the failure form of coal rock mass is triangle block shear slip. Gao et al. (2014) stated that slip failure is a common phenomenon in underground coal mine roadways, which is often induced by the high horizontal stress. With the transition of crustal stress and the extension of shear fractures in coal rock mass, slip failure may suddenly occur. ...
... He et al. (2018) investigated the influences of spatial configuration of discontinuities on an unsupported rock tunnel. With UDEC Trigon approach, Gao et al. (2014) simulated shear failure in coal mine roadways. Zheng et al. (2018) studied the mechanisms of flexural toppling failure in anti-inclined rock slopes. ...
... In Fig. 9, there are shear bans and unstable triangular blocks in the coal rock mass whether or not there are weak interlayers in coal rock masses, which indicates that the unstable triangular blocks are the common potential disaster triggers in underground coal mining. Furthermore, it can be found that the unstable triangular blocks are most caused by shear failure, which is a common failure mechanism in underground coal mine roadways (Gao et al. 2014). ...
To study the characteristics and mechanism of slip failure in coal rock mass containing weak interlayers in deep underground, UDEC was used to investigate coal rock mass. The results show that the weak interlayer can change the path of stress transfer, and with the increase of thickness of the weak interlayer, the horizontal displacement and stress all show an increasing trend, and the closer the monitoring point is to the wall of roadway, the greater the horizontal displacement. The depth of plastic zone also increases with the growth of the thickness of weak interlayer. Additionally, the lateral pressure coefficient has a similar influence on the horizontal displacement and stress with the thickness of weak thickness. The horizontal and vertical stresses of the coal rock mass with no weak interlayers are often the largest before the peak stress and the smallest after the peak stress. These results can help to understand the formation mechanism of slip failure in coal rock mass in deep underground and is of great significance for evaluating the safety and taking reasonable supporting measures.
... Một số cấu trúc hạt điển hình có thể kể tên như Voronoi 9 hay Trigon 10 (Hình 2). Một số nghiên cứu đã áp dụng thành công mô hình nền tảng hạt trong mô phỏng DEM để nghiên cứu sự ổn định khối đá ở phạm vi thực địa như bờ dốc mỏ 12 , đường lò mỏ [13][14][15] , đá vách lò chợ 16,17 , sụt lún bề mặt địa hình mỏ 18 , gương lò chợ 19 , và trụ than bảo vệ đường lò mỏ 20 . Mặc dù các nghiên cứu trên đã góp phần làm sáng tỏ cơ chế ổn định của khối đá trong các điều kiện khai thác mỏ khác nhau, việc áp dụng mở rộng các kỹ thuật mô phỏng tương ứng là khó khăn do chưa thống nhất: (i) cách thức lựa chọn kích thước khối hạt; (ii) quy trình hiệu chỉnh tính chất vật liệu (xem Mục 2); và (iii) trình tự mô phỏng thí nghiệm và giám sát. ...
... Giá trị góc nội ma sát và dính kết được chọn lần lượt là 1 MPa và 8 độ để độ bền mẫu mô phỏng là 4.44 MPa (Mô phỏng14). Trạng thái mẫu sau phá hủy và biểu đồ ứng suất -biến dạng thể hiện trong Hình 7. Biểu đồ tương quan biến dạng dọc trục -biến dạng hông thể hiện trong Hình 8. 4: Tính chất cơ học mô hình UDEC-GBM trong quá trình hiệu chỉnh Giá trị độ bền kéo mẫu đá trong Mô phỏng 14 thu được là 0.35 MPa. ...
Sự ổn định khối đá xung quanh lò chợ hoặc công trình ngầm trong mỏ than hầm lò là điều kiện tiên quyết để hoạt động sản xuất than diễn ra bình thường và đảm bảo an toàn. Tuy nhiên, các lò chợ thực tế vẫn thường xảy ra các sự cố mất ổn định tại gương khai thác hoặc trên nóc gây tai nạn lao động nghiêm trọng, thậm chí chết người. Các nghiên cứu về sự mất ổn định này cho tới nay chưa làm rõ thỏa đáng sự hình thành và phát triển phá hủy khối đá trong biểu hiện, nguyên nhân phần lớn là bởi đặc trưng bất đẳng hướng, không liên tục, không đồng nhất và không đàn hồi của khối đá mỏ. Nhằm phục vụ nghiên cứu biểu hiện ổn định khối đá nêu trên, nội dung bài báo trình bày một nghiên cứu hoàn thiện kỹ thuật mô phỏng số khối đá xung quanh lò chợ bằng mô hình nền tảng hạt. Kỹ thuật mô phỏng sử dụng phương pháp phần tử rời rạc và cấu trúc hạt Voronoi. Một quy trình hiệu chỉnh tính chất vật liệu trong kỹ thuật mô phỏng được đề xuất và kiểm chứng tính đúng đắn thông qua áp dụng cho đất đá thực địa mỏ than Hà Lầm, tỉnh Quảng Ninh. Nội dung bài báo cũng trình bày một kỹ thuật lựa chọn kích thước khối hạt đa giác và giám sát mô phỏng hiệu quả. Quy trình đề xuất cùng với bộ mã lập trình mô phỏng tương ứng sẽ phục vụ đắc lực công tác nghiên cứu ổn định khối đá xung quanh công trình ngầm, từ đó đề xuất các giải pháp kỹ thuật đảm bảo an toàn và sản xuất theo kế hoạch.
(http://stdjsee.scienceandtechnology.com.vn/index.php/stdjsee/article/view/711)
... Among the above, the occurrence mechanisms should be studied clearly first, to form the basis of follow-up monitoring, early warning and treatment measures. There are many studies on the mechanism of rock bursts, which mainly include energy theories [13], strength theories [14], stiffness theories [15], rock burst tendency theories [16], instability theories [17], three factor theories [18], three criteria theories [19], dynamic and static load principles [20], rock burst initiation theories [21], butterfly rock burst mechanism [22,23], etc. No matter which mechanism, it must be aimed at the rock mass involved in the disaster, the mechanical properties of the rock mass, the stress environment leading to the disaster, the time process of the disaster, etc. ...
... In terms of rock burst research methods, some mechanism studies aimed at a specific condition, such as fault [24,25], strong mining influence [26,27], high stress environment [28], high dynamic load [29], stratum movement [30], etc. Others study a series of problems from one aspect, such as strength [31], stress [32], energy [33], gas superposition [34], surrounding rock properties [35][36][37] and so on [18]. Whether it is into specific conditions or a certain aspect, the research plays a positive role in revealing the rock burst mechanism. ...
It is very necessary to study the mechanism of rock burst, which is related to the safe construction of many geotechnical projects. Previous studies have shown that small trigger stress will lead to large energy release, but the specific conditions that cause the release and how to quantify the energy are urgent problems to be solved. In this study, an innovative calculation method of rock mass energy release is proposed, and the calculated release energy is consistent with the monitoring results of field monitoring equipment. The revealed mechanism of rock burst reflected is that under the condition of a large-ratio pre-state stress field (mostly > 2.5), a small trigger stress field will lead to a large amount of energy release under “late butterfly shape” or “final butterfly shape” of the plastic zone. This study reveals the key factor of rock burst, which plays an important reference role for the mechanism research, subsequent monitoring and treatment method of rock burst.
... The roof rock strata structures mainly included upper-soft and lowerhard strata, upper-hard and lower-soft strata, containing faults strata, and regenerated roof strata as well as laminated-jointed roof strata [6][7][8][9][10][11]. Above research results showed that the characteristics of roof strata structure have a significant impact on the roof instability mode of roadway, especially for large-span roadway. Gao et al. [12] investigated the shear failure phenomenon of a shale roadway roof by using an innovative numerical approach and confirmed the time sequence of marked microseismic activity, significant stress changes, and accelerated displacement during the process of a roof fall. Shen et al. [13] monitored the displacement, stress changes, and seismic activities during a roof fall by using an integrated roof monitoring system and found that the changes associated with roof falls follow a time sequence of initial seismicity, followed by stress changes and then displacement. ...
... Gao et al. [12] studied the failure mode of roadway roof without the discontinuous surface, and they believed that shear failure of the roadway roof initiates at the roadway corners and then progressively propagates deeper into the 12 Geofluids roof, finally forming a large scale roof failure. But, the roof failure model is symmetrical as shown in Figure 17(a). ...
Existence of roof discontinuity surface is one of the extremely important factors causing the asymmetric fracture of roadway roof, especially for large-span soft rock roadway. In this paper, a crack density coefficient (D) is firstly defined by local thresholding-microcell segmentation method and used to analyze the evolution law of D with roof drilling depth (d) (the lower and upper of roof discontinuity are defined as regions I and II, respectively). Then, UDEC numerical simulation is used to investigate the asymmetry evolution law of roof total displacement, maximum principal stress, and crack density with stress release coefficient (α) considering the effect of discontinuity surface. Research results indicate that (1) the roof parameters (D) in regions I and II both show a negative logarithmic function decreasing trend with the increase of drilling depth. When d
... Shi et al. [10] investigated the crack evolution mechanism of the gob-side entry for different conditions and proposed optimized-support parameters combined with roof-cutting measures. Gao [11][12][13] carried out a series of numerical simulations using the UDEC Trigon approach to focus on the roadway damage caused by squeezing failure and shear failure, and the effects of rock bolts in the roadway support were also evaluated. FLAC3D can simulate the mechanical behavior of geological materials and geotechnical engineering effectively and is widely used in underground coal mining activities [14][15][16]. ...
... Combining (9)- (12) and substituting the data, the breaking position of main roof is 4.4~5.8 m from the edge of the gob. ...
In multi-seam mining, the residual coal pillar (RCP) in the upper gob has an important influence on the layout of the roadway in the lower coal seam. At present, few papers have studied the characteristics of the surrounding rock of gob-side entry driving (GED) with different coal pillar widths under the influence of RCP. This research contributes to improving the recovery rate of the extra-thick coal seam under this condition. The main research contents were as follows: (1) The mechanical parameters of the rock and coal mass were obtained using laboratory experiments coupled with Roclab software. These parameters were substituted into the established main roof structure mechanics model to derive the breakage position of the main roof with the influence of RCP, and the rationality of the calculation results was verified by borehole-scoping. (2) Based on numerical simulation, the evolution laws of the lateral abutment stress in the lower working face at different relative distances to the RCP were studied. FLAC3D was used to study the whole space-time evolution law of deviatoric stress and plastic zone of GED during driving and retreating periods with various coal pillar widths under the influence of RCP. (3) The plasticization factor P was introduced to quantify the evolution of the plastic zone in different subdivisions of the roadway surrounding rock, so as to better evaluate the bearing performance of the surrounding rock, which enabled a more effective determination of the reasonable coal pillar width. The field application results showed that it was feasible to set up the gob-side entry with an 8 m coal pillar below the RCP. The targeted support techniques with an 8 m coal pillar could effectively control the surrounding rock deformation.
... Empirically, Esterhuizen [5] stated that the laminated roof fallcaused cavities can have near-vertical sides, as opposed to dome-like failure cavities formed in rocks that are not bedded, which is exactly the scenario observed in our entry models. Numerically, our model results with the intact roof were consistent with that obtained with a Trigon bonded block model (BBM) [31] shown in Fig. 15a, verifying the capacity of our inputs of block material. Moreover, the laminated roof failure observed by Voronoi BBM [32] presented similar failure modes observed in our models shown in Fig. 7c, verifying the combination of block material and discontinuities' parameters. ...
... Roof failure model results using a intact roof simulated with Trigon BBM[31] and b laminated roof simulated with Voronoi BBM[32] ...
The effect of discontinuities on the fracturing and mechanical behavior of shale has been extensively investigated on a laboratory scale in previous works. It is well agreed that the lamination properties, including discontinuity and lamina properties, affect the behavior of shale. However, it is still unclear how the lamination properties are affecting the stability of the shale roof in an underground coal mine entry. This paper investigated the effect of lamination properties using discrete element method on a mine-scale entry model as an extension to the previous work conducted on laboratory scale models (Q. Shi and B. Mishra, Discrete Element Modeling of Delamination in Laboratory Scale Laminated Rock, Mining, Metallurgy & Exploration, vol. 37, no. 5, pp. 1–14, Sep. 2020). The microparameters for both the laminas and discontinuities were calibrated with laboratory data. In the calibration, a numerical laminated Brazilian disc was created and tested for comparison with laboratory results. The effects of lamina thickness, discontinuity strength, and supporting pressure on the model’s roof strength and the stress distribution were also investigated. Numerical results showed that the lamination properties and supporting pressure contribute significantly to the stress distribution in the roof and its stability. The horizontal stress at a fixed depth in the roof increased with the lamina thickness, discontinuity strength, and supporting pressure. The laminated roof strength was found to increase with the increase of lamina thickness but never exceeds the strength of an intact roof comprising the only matrix.
... Empirically, Esterhuizen [5] stated that the laminated roof fallcaused cavities can have near-vertical sides, as opposed to dome-like failure cavities formed in rocks that are not bedded, which is exactly the scenario observed in our entry models. Numerically, our model results with the intact roof were consistent with that obtained with a Trigon bonded block model (BBM) [31] shown in Fig. 15a, verifying the capacity of our inputs of block material. Moreover, the laminated roof failure observed by Voronoi BBM [32] presented similar failure modes observed in our models shown in Fig. 7c, verifying the combination of block material and discontinuities' parameters. ...
... Roof failure model results using a intact roof simulated with Trigon BBM[31] and b laminated roof simulated with Voronoi BBM[32] ...
The effect of discontinuities on the fracturing and mechanical behavior of shale has been extensively investigated on a laboratory scale in previous works. It is well agreed that the lamination properties, including discontinuity and lamina properties, affect the behavior of shale. However, it is still unclear how the lamination properties are affecting the stability of the shale roof in an underground coal mine entry. This paper investigated the effect of lamination properties using discrete element method on a mine-scale entry model as an extension to the previous work conducted on laboratory scale models (Q. Shi and B. Mishra, Discrete Element Modeling of Delamination in Laboratory Scale Laminated Rock, Mining, Metallurgy & Exploration, vol. 37, no. 5, pp. 1–14, Sep. 2020). The microparameters for both the laminas and discontinuities were calibrated with laboratory data. In the calibration, a numerical laminated Brazilian disc was created and tested for comparison with laboratory results. The effects of lamina thickness, discontinuity strength, and supporting pressure on the model’s roof strength and the stress distribution were also investigated. Numerical results showed that the lamination properties and supporting pressure contribute significantly to the stress distribution in the roof and its stability. The horizontal stress at a fixed depth in the roof increased with the lamina thickness, discontinuity strength, and supporting pressure. The laminated roof strength was found to increase with the increase of lamina thickness but never exceeds the strength of an intact roof comprising the only matrix.
... Bai et al. [27] used UDEC Voronoi method to study the progressive failure process of roadway roof of large section, and the results showed that shear cracks dominated the roadway roof, and they put forward reasonable control technology. Gao et al. [28] used UDEC Trigon model to simulate the shear failure process of the roof in large section and successfully captured the shear failure of roof characterized by crack initiation and propagation. e results showed that the shear failure of roadway roof started at the corner of roadway, then gradually spread to the depth of roof, and finally formed a large-scale roof failure. ...
... In the trigon method, rock mass is expressed as a combination of triangular blocks bonded by internal contact to simulate brittle materials [28,34]. It is assumed that each triangular block is an elastic material and is divided into triangular finite difference domains, which cannot fail. ...
The stability of large section open-off cut in deep mines (LODM) is the key factor affecting the normal equipment installation and safe mining in fully mechanized top-coal caving face. The mechanical model shows that the deflection of the roof of the LODM is proportional to the cubic of span. In this paper, UDEC Trigon model is established, and the parameters of different coal measures strata are modified in detail. The evolution law, failure mode, and damage degree of roof cracks in secondary tunneling are studied, and the roof support effect is analyzed. The numerical simulation results show that the process of roof crack evolution after the primary excavation section and the second excavation section can be divided into three stages according to microseismic activities, and the reasonable supporting time can control the propagation of roof microcracks and reduce the development height of macrocracks. The rock bridge existing in the roof rock stratum after the combined support of long and short anchor cables can effectively limit the formation of macrocracks and their interaction; especially the key support in the interface area can reduce the development height of roof cracks in secondary tunneling and weaken the damage degree of roof rock stratum in the LODM. The field test shows that the moved volume of rib-to-rib and roof-to-floor of the LODM is stable at about 350 mm and 550 mm, respectively. The numerical simulation in this paper is helpful to understand the failure mode of roof in LODM with large mining height and provides a method for the design of its control technologies.
... In terms of numerical investigations of rock-support interaction, there is only a limited number of studies that have been conducted to date. 33,34,46,47,52 Gao et al., 52 Kang et al. 33 and Yang et al. 46,47 presented a comparison of roof deflection with and without support in a coal mine entry, while Bai et al.34 presented some qualitative support guidelines for yielding coal pillars based on a discontinuum model (supports were not tested in the models). As a first step towards understanding the influence of reinforcement on ground behavior in numerical models, Sinha and Walton 23 employed the elastic W/H = 2 BBM from Sinha and Walton 38 to evaluate the effect of reinforcement on pillar behavior. ...
... In terms of numerical investigations of rock-support interaction, there is only a limited number of studies that have been conducted to date. 33,34,46,47,52 Gao et al., 52 Kang et al. 33 and Yang et al. 46,47 presented a comparison of roof deflection with and without support in a coal mine entry, while Bai et al.34 presented some qualitative support guidelines for yielding coal pillars based on a discontinuum model (supports were not tested in the models). As a first step towards understanding the influence of reinforcement on ground behavior in numerical models, Sinha and Walton 23 employed the elastic W/H = 2 BBM from Sinha and Walton 38 to evaluate the effect of reinforcement on pillar behavior. ...
In this study, Bonded Block Models (BBMs) are used to investigate the pillar damage mechanisms and rock-support interaction in massive-to-sparsely-fractured rockmasses. Hypothetical granite pillar models of width-to-height (W/H) ratio of 1, 2 and 3 are developed, and the input parameters are constrained by matching the stress-strain response of the BBMs to the stress-strain curves from FLAC3D models that were
previously calibrated to an empirical pillar strength database. Two different block representations are also considered – elastic and inelastic. It was found that inelastic blocks are necessary to capture the behavioral transition from strain-softening to pseudo-ductile with increase in pillar W/H. Post-calibration, different rockbolt combinations are tested in the BBM and their influence on the pillar strength and lateral deformations are analyzed. It was found that as the support density is increased, the peak pillar strengths
also increase but the effect is dependent on the W/H. Deformation of the outer stress-fractured region and bulking systematically decreased with increasing support density, but the exact trend evolved as the pillars were loaded to various points on their stress-strain curves. Lastly, a BBM pillar was developed with explicit intra-block fracturing capability (i.e., individual blocks could break) and the support analysis was repeated.
The goal was to understand if the continuum representation of damage within the inelastic blocks led to some underestimation of the rock-support interaction mechanism. It was ultimately concluded that the continuum inelastic representation of smaller-scale damage within individual blocks allows for a more appropriate representation of the rock-support interaction than the explicit intra-block representation.
... The initial contact normal stiffness K n and shear stiffness K S are calculated through calculation. However, in order to obtain the most precise properties feasible, a sequence of uniaxial compression tests has to be conducted in the UDEC software [35][36][37]. Through ongoing modification of the contact properties in UDEC, Young's modulus and compressive strength of the numerical sample exhibit congruence with those observed in the laboratory. ...
... The results show that the distribution of energy release density and plastic zone has a strong management, and has experienced a process from small butterfly to large butterfly. From the value of energy distribution, the 20 The calculation results of the released energy of rock mass are shown in Figure 6, from which it can be seen that the phased characteristics are obvious. The entire process shown in the figure can be divided into three stages: the pregnant period, growth period, and upheaval period, representing the phenomenon of releasing energy from small to large, respectively, and then rapidly increasing. ...
The research on the formation factors of rock burst is one of the main research directions of rock mechanics in recent years, which is helpful to solve the problem of rock burst accidents. So, in this study, the calculation method of energy released during rock burst is first obtained by using different medium models, and then, the formation factors of rock bursts are obtained by comparing the calculation energy with the actual accident energy. The method of energy calculation utilizes the difference between elastoplastic and pure elastic models to innovatively quantify the specific values of energy released before and after the occurrence of the rock burst. It is considered that the stress and plastic zone state before the occurrence of rock burst have an important influence on the occurrence of the accident and are one of the formation factors, while the deviatoric stress field and butterfly-shaped plastic zone create conditions for greater energy release. In addition, the trigger stress constitutes another formation factor. The plastic zone state before rock failure is verified by the experimental test; the location distribution shape of acoustic emission (AE) events during the later stage of compression failure is approximately the same as theoretical result. The results also preliminarily indicated the fractal characteristics of acoustic emission events distribution before sample failure. The study obtained the formative factors of rock burst accident, which provides a new ideas and references for the research on the formation of rock bursts.
... Previous studies by the authors have demonstrated the realism of rock-support interaction behaviors predicted by BBMs [27,37]. There are also numerous numerical [73][74][75][76] and empirical [6,77] studies in the literature that have findings that are consistent with significant drops in the rib and roof deformations due to the installation of rockbolts. ...
Discontinuum Bonded Block Modeling (BBM) represents a potential tool for support design, as these models can reproduce both the rock fracturing process and the influence of reinforcement on unsupported ground. Despite their strengths, discontinuum models are seldom used for mining design due to their computationally intensive nature. This study is an application of an integrated 3D continuum-2D discontinuum approach, in which the mine-wide stress distribution process is modeled using a continuum software, and the local deformation behavior in response to a strain path from the continuum model is simulated with a 2D discontinuum software. In June 2017, two multi-point borehole extensometers were installed in a longwall chain pillar to record ground displacements as a function of the longwall face position. The data from one of the extensometers were employed to calibrate a panel-scale FLAC3D model. The boundary conditions along the pillar slice containing the extensometer were extracted from the FLAC3D model and applied to a 2D BBM, and the input parameters were modified to match the extensometer data. The calibrated BBM was able to reproduce the unsupported rib deformation and depth of the fracturing well. Subsequently, a few support schemes were tested to demonstrate how the incorporation of support might affect rib deformation.
... Due to the complex deformation characteristics of the surrounding rock in inclined coal seam roadways, anchor cables play a crucial role in the design of support systems for these roadways. However, studies [20][21][22] show that when the roadway is driven, the surrounding rock will shift and slip around the joint surface; at the coal roadway support site, tensed shear failure occurs in some anchor cables, mostly in the free section, affecting the stability of the roadway. There are many research studies on the shear performance of anchor cables at home and abroad. ...
In view of the complex asymmetric deformation characteristics of inclined coal seam roadways and the tensed shear failure of anchor cable supports, the asymmetric support scheme of an anchor cable with a C-shaped tube is proposed. In order to study its supporting effect on an inclined coal seam roadway, this paper first explores the difference in shear performance between an anchor cable with a C-shaped tube and an anchor cable through double shear tests. Then, based on the asymmetric deformation characteristics of an inclined coal seam roadway in the Pangpangta Mine, a numerical simulation is used to study the asymmetric support effect of an anchor cable with a C-shaped tube in an inclined coal seam roadway. The results of the double shear test show that the anchor cable with the C-shaped tube has stronger resistance to shear load than that of the anchor cable. Through the results of the numerical simulation, the original stress field distribution on both sides of the roadway was found to be uneven due to the influence of the coal seam dip angle, and after the excavation of the inclined coal seam roadway, the displacement and plastic zone distribution on both sides showed obvious asymmetric characteristics. Compared with the symmetric support, the asymmetric support can obviously alleviate the asymmetric deformation characteristics of the two sides and effectively control the deformation and plastic failure zone of the roadway. The anchor cable with the C-shaped tube has better resistance to shear deformation than that of the anchor cable. The anchor cable with the C-shaped tube can reduce the deformation and plastic area of the roadway more effectively.
... Through the UDEC (a discrete element software), many scholars can intuitively see the fracture development and damage degree in the coal pillar of gob-side entry driving [78][79][80][81]. By analyzing the crack propagation morphology and plastic state in the coal pillar, they can guide the reasonable setting position of the gob-side entry and coal pillar width to minimize the influence of the coal pillar on the stability of the roadway [82][83][84]. ...
China has abundant coal resources, and the distribution of coal seams is complex. Thick coal seams account for more than 45% of all coal seams. Fully mechanized top coal caving mining has the advantages of large production, high efficiency, and low cost. In fully mechanized caving mining, especially in fully mechanized caving mining of extra-thick coal seams, the mining space is ample, the mine pressure is severe, and the roadway maintenance is complex. As a result, it is necessary to summarize and discuss the gob-side entry driving of fully mechanized caving in theory and technology, which will help to promote the further development of fully mechanized caving gob-side entry driving technology. First, in recent years, the research hotspots of gob-side entry driving have focused on the deformation mechanism and the control method of the roadway surrounding rock. Secondly, this paper discusses the theoretical models of the “triangle-block” and “beam” for the activity law of the overlying strata in gob-side entry driving, including the lateral breaking “large structure” model, compound key triangle block structure model in the middle and low position, the high and low right angle key block stability mechanics model, elastic foundation beam model, low-level combined cantilever beam + high-level multilayer masonry beam structure model, and the vertical triangular slip zone structure model. It introduces the “internal and external stress field theory” and the “stress limit equilibrium zone model”. Thirdly, it summarizes several numerical simulation analysis methods in different conditions or research focuses and selects appropriate constitutive models and simulation software. Finally, it introduces surrounding rock control technology, including two ribs, the roof, and under challenging conditions. It provides a method reference for support in similar projects.
... The layered rock mass is used as a laterally isotropic material to explain the anisotropic behavior in tunneling (Tien et al., 2006;Fortsakis et al., 2012). Further research has been conducted on the deformation and shear failure mechanism of soft and hard rock interbedding (Meng et al., 2013;Gao et al., 2014;Yang et al., 2018;Li et al., 2020;Zhou and Yang, 2021). The standard deformation classification methods are the Jethwa method (Jethwa et al., 1980), the relative strain method (Anagnostou, 1993), the Aydan method, the comprehensive coefficient method (Panthi and Nilsen, 2007) and the Hoek method (Hoek and Guevara, 2009). ...
Due to the influence of composite strata, traditional monitoring and measurement cannot accurately reflect the nonuniform deformation and corresponding local damage to tunnels. However, the nonuniform deformation precise analysis and effective prediction of composite strata tunnels are indispensable in related projects. This analysis established a nonuniform deformation classification method and 135 groups of numerical models for a composite strata tunnel. The nonuniform deformation classification method involves four reference-defined indexes , the average deformation coefficient, variable deformation coefficient, segmental deformation coefficient, and abnormal deformation coefficient. The numerical model took the dip angle of the formation, formation proportion, in-situ stress, and pore pressure as composite strata model conditions, with four reference indexes in the nonuniform deformation classification method as the output parameters. The intelligent model of the least squares supports vector machine (PSO-LSSVM) was innovatively utilized to predict nonuniform deformation in composite strata during tunnel excavation. The prediction results of the PSO-LSSVM are consistent with actual engineering data relying on the engineering situation of Teermo Tunnel. This approach can provide targeted guidance for a composite stratum tunnel engineering evaluation system.
... For the roadway problems of a coal mine, scholars around the world have made many efforts to study the properties of the roadway surrounding rock, the state of the deposit, the influence of mining stress, the mechanism of the surrounding rock deformation, and the role of support, etc., and have accumulated rich theoretical and practical engineering experience. Gao et al. [9][10][11] conducted a series of numerical simulations to study the roadway's squeezing failure and shear failure, using the UDEC Trigon method and analyzed the effects of rock bolts on the roadway support. Wang et al. [12] tested the properties of surrounding rock considering the influence of the lower coal seam on the stress state by a numerical simulation. ...
A large coal pillar (usually more than 90 m) is generally left in place to ensure the stability of main roadway groups, due to its long service lifespan, which commonly also causes a significant loss of coal resources. The design of the width of the protective coal pillar and the control system of the surrounding rock are directly determined by the characteristics of the stress field and the damage mechanism under the influence of the mining activities. However, there are few studies on the effects of the partial-stress boosting and the direction deflection of the stress field on the failure evolution of the surrounding rock (especially in multi-seam mining). In this paper, theoretical analysis and numerical simulation are used to investigate the direction evolution of the maximum principal stress in front of the working face with malposition distances between the upper and lower working faces during the influence of double coal seams mining. Furthermore, a large-scale numerical model is used to study the deviatoric stress evolution of the surrounding rock and the propagation process of the plastic zone in the main roadway group with different widths of protective coal pillars. Then, an asymmetric cooperative anchoring classification method is proposed to strengthen the roadway support, depending on the critical area of the deviatoric stress in the roadway surrounding rock. The peak zone deflection of the deviatoric stress determines the evolution direction of the plastic area, and the peak value of the deviatoric stress presents a typical asymmetric stress boosting on both sides of the roadway. These findings are validated by the on-site ground pressure monitoring results and the practical failure modes of the surrounding rock.
... In the calculation, the elastic constitutive model was adopted for the rock block, and the coulomb slip constitutive model was adopted for the contact surface. It was assumed that the mechanical properties of the block contact conformed to the linear stress-strain relationship [8], as shown in Fig. 13. ...
Underground engineering excavation can lead to sharp stress change in the rock mass around the excavation surface, which can cause different degrees of rock damage, ultimately resulting in instability failure. Especially for inclined stratified rock mass that is ubiquitous on Earth, the evolution characteristics, development law and formation mechanism of an excavation damage zone are highly complicated due to its significant asymmetry. Therefore, the evolution mechanism and deformation failure properties of a typical deep roadway in inclined rock strata in Jinchuan Mine of China were investigated by means of a field investigation, theoretical analysis, similar model test and numerical simulation. The results indicate that the deformation failure of a roadway in deep inclined rock strata shows a prominent asymmetry and time sequence. Ground stress has a great influence on the development mode and evolution characteristics of the surrounding rock damage zone. However, as a deep ground stress environment tends to cause hydrostatic pressure, its leading role is gradually weakened. The structural planes control the damage evolution mode of the surrounding rock, an excavation damage zone developed parallel to the interface is formed around the goaf, and an overall instability of the roadway is caused by the sliding of surrounding rock along the structural plane. The conclusions of this study should provide a theoretical reference and demonstrate the key technologies that support underground engineering under similar geological conditions.
... The Mohr-Coulomb criterion was used in the whole model, and the model parameters are shown in Table1. The left and right boundary of the model denoted limited horizontal displacement (Gao et al., 2014;Zhang et al., 2020b;Vergara et al., 2020;Wu et al., 2021;Kim and Larson, 2022). The lower boundary was limited to vertical displacement. ...
We determine the key issues of the reasonable supporting capacity of a shield and the shield selection in the panel face in a deep and thick coal seam. The No.232204 panel face of the Meihuajing mine in the Yuanyanghu mining area is taken as the research background. Using theoretical analysis, numerical simulation, and field measurement, the fracture characteristics and instability forms of roof structure of the panel face were studied. A mechanical model of roof fracture structure was established to calculate the reasonable supporting capacity of the shield, which guided the shield selection for the panel face. Also, a calculation and analysis system of the supporting capacity of the shield in a deep and thick coal seam was developed to realize the dynamic calculation and analysis of the supporting capacity of the shield. The results are as follows. 1) when the first weighting of the main roof appears, the broken rock block is hinged to form a “three-hinged arch” balanced structure. When the periodic weighting of the main roof appears, the broken rock block forms a “voussoir beam” balance structure. Also, the supporting capacity of the shield is stable between 8,900 and 9,600 kN. 2) The theoretical calculation showed that the supporting capacity of the shield in the No.232204 panel face was 9,581.04 kN, and the ZY10000/28/62D shield with supporting capacity of 10,000 kN is selected in the No.232204 panel face. 3) Through self-developed calculation software, the thickness and bulk density of the immediate roof had little influence on the supporting capacity of the shield, and the main roof thickness and bulk density have considerable influence on the supporting capacity of the shield. 4) The field measurement showed that the maximum supporting capacity of the shield in the panel face was distributed between 9,000 and 10,000 kN, which accounted for 77.58%. The loading utilization rate was more than 90%, and therefore the selected shield was reasonable. The research results provide a theoretical basis for the selection of shield supports for a deep buried and thick coal seam in the Yuanyanghu mining area.
... In the calculation, the elastic constitutive model was adopted for the rock block, and the coulomb slip constitutive model was adopted for the contact surface. It was assumed that the mechanical properties of the block contact conformed to the linear stress-strain relationship [8], as shown in Fig. 13. ...
... Numerical methods provide an effective approach to capture these features. e numerical methods can be classified into continuum methods (such as boundary element method (BEM), finite element method (FEM), finite-difference method (FDM)), and discontinuum methods (such as distinct element method (DEM) and discontinuous deformation analysis (DDA)) and hybrid methods (such as FEM/DEM and FDM/ DEM) [17,18]. Continuum methods have been used to simulate longwall mining for decades but have impactful limitations. ...
It is usually difficult to capture strata caving and gob evolution characteristic in longwall mining at engineering scales. This paper uses bonded block modelling (BBM) approach within a distinct element method (DEM) code to simulate strata behaviour in longwall mining, which captures the caving phenomena and bulking characteristics of roof strata successfully. Many features in longwall mining, including the caving and compaction of gob strata and the associated stress evolution, are reproduced in the model. Four zones in longwall gob are identified based on its stress characteristics: voussoir influencing zone, compacted zone, compacting zone, and pilling zone. The initial bulking factor of the caved strata ranges from 1.12 to 1.25 and decreases gradually to the residual bulk factor of approximately 1.05 as the longwall face advances. The caved strata in the longwall gob present strain hardening behaviour and the load carrying capacity increases exponentially as a function of strain. Moreover, the range of the interaction between the caving strata and the overburden in gob was discussed, which provides a reference when using a continuum method to simulate longwall mining.
... The destruction of the soil, as well as the roof rocks begins with stratifi cation of bands, layers. Stratifi cation is one of the most characteristic features of joint deformation and failure of a stratum of layered rocks [10]. The places of rock separation are determined by the ratios of their strength and deformation properties, the magnitude of normal and shear stresses, the magnitude of the bond forces of the layers along the contact surfaces. ...
Purpose. Development of a methodology for assessing the occurrence of sudden fractures in the soil on the example of the development of seam d6 of the Karaganda coal basin. Methodology. Theoretical and experimental studies on the gas content of a coal seam, analysis of technologies for its development. Findings. Studies have shown that sudden fractures of the overworked massif with and without a breakthrough of methane are of the same nature. A necessary condition for the formation of sudden fractures of the seam soil (SFSS) in a mine is the presence of a pack of coal or rock with reduced strength and high gas content in the soil; in this case, the breakthrough of methane into the development occurs only if the cracks formed in the soil of the workings reach the sources of gas release gas-bearing layers (interlayers) of coal. Originality. A predictive indicator of the hazard of the occurrence of SFSS in the working face is proposed, a dimensionless criterion which is the product of dimensionless coefficients reflecting the influence of the development depth, gas content of the seam, the width of the bottomhole zone, the thickness of the protective layer, the thickness of the crumpled coal pack of the lower layer, longwall withdrawal from the assembly chamber, the length of the working face. A nomogram is developed for assessing the risk of SFSS in the development of reservoir d6. Practical value. According to the results of the research, it was proposed to predict the SFSS at the design stage, for the early implementation of regional measures for their prevention. To assess the WSP, a nomogram is to be used for the prevention of SFSS, which includes advanced development of protective seams, measures to reduce the ability to delaminate coal in the lower layer, change in technology for coal mining with intensive degassing of the developed seam and use of interval hydraulic fracturing.
... The disaster accident of roadway surrounding rock is the result of the interaction between the rock body and the coal body [1][2][3]. If only studying the physical and mechanical properties of coal body cannot effectively guarantee the safety of coal mine production, it is necessary to focus on the physical and mechanical properties of the "roof-coal" combination coal and rock mass [4][5][6][7][8][9][10]. At the same time, water injection can effectively reduce the probability of rock burst occurrence in rock burst mines [11][12][13][14][15][16][17]. ...
Coal and rock are often in an environment of hydraulic-mechanical coupling. In order to study the failure mechanism of the coal-rock combined body under the coupling action of hydraulic and mechanical, the RFPA-Flow software was used to analyze the failure mode, strength change law, and acoustic emission change law of the coal-rock combined body with different rock-to-coal height ratios under the combined action of uniaxial load and water pressure. The research results show that the peak strength, residual strength, and stress drop of coal-rock combined body with different rock-to-coal height ratios decrease after water pressure, which reduces the occurrence probability of rock bursts. However, the stress showed a vertical drop phenomenon in the later stage of loading, indicating that the coal-rock combined body still maintains the characteristics of brittle failure after being softened by water, and the roadway may still have rock bursts. The research conclusions can provide a theoretical basis for using water injection measures to prevent rock bursts in deep coal and rock masses.
... It was observed that the increasingly complicated engineering geological conditions had posed a great threat to the safety production of Chinese coal mines. Considerable efforts have been expended to acquire a comprehensive understanding of mechanism of the dynamic failure induced by different factors, including roof rupture [38][39][40], fault slip [8,26,41], fold structure [42,43], coal pillar [7,44,45], etc. However, more complex geotechnical conditions are characterized by composite geological structures that have been threatening the safety in underground mining. ...
With the increase in coal mining depth, engineering geological conditions and the stress environment become more complex. Many rock bursts triggered by two combined faults have been observed in China, but the mechanism is not understood clearly. The focus of this research aims at investigating the influence of two combined faults on rock burst mechanisms. The six types of two combined faults were first introduced, and two cases were utilized to show the effects of two combined faults types on coal mining. The mechanical response of the numerical model with or without combined faults was compared, and a conceptual model was set up to explain the rock burst mechanism triggered by two combined faults. The influence of fault throw, dip, fault pillar width, and mining height on rock burst potential was analyzed. The main control factors of rock burst in six models that combined two faults were identified by an orthogonal experiment. Results show that six combinations of two faults can be identified, including stair-stepping fault, imbricate fault, gra-ben fault, horst fault, back thrust fault, and ramp fault. The particular roof structure near the two combined faults mining preventing longwall face lateral abutment pressure from transferring to deep rock mass leads to stress concentration near the fault areas. Otherwise, a special roof structure causing the lower system stiffness of mining gives rise to the easier gathering of elastic energy in the coal pillars, which makes it easier to trigger a rock burst. There is a nonlinear relationship between fault parameters and static or dynamic load for graben faults mining. The longwall face has the highest rock burst risk when the fault throw is between 6 and 8 m, the fault dip is larger than 65°, the mining height is greater than 6 m, and the coal pillar width is less than 50 m. The stair-stepping, imbricate, horst, and ramp fault compared to the other fault types will produce higher dynamic load stress during longwall retreat. Fault pillar width is the most significant factor for different two combined faults, leading to the rise of static load stress and dynamic proneness.
... Since 2010, a few standard calibration processes were established for the derivation of a unique set of contact micro -properties for matching on -site rock mass response (Kazerani and Zhao, 2010;Kazerani, 2013;Gao and Stead, 2014). Due to the standard process in calibration and advantage in modelling failure development, GBM has been widely used in UDEC for studying longwall stability such as roadway (Coggan et al., 2012;Gao et al., 2014c;Bai et al., 2016b), roof strata (Gao et al., 2014a;Li et al., 2016), surface subsidence (Zhang et al., 2017), coal face (Yao et al., 2017), and coal pillar (Wu et al., 2019) (Figures 6 -7). The use of GBM, however, must take into consideration the disadvantages and uncertainties of the technique. ...
Longwall-associated geotechnical problems have been extensively studied by using numerical modelling methods. However, proper representation of its geological structures remains a challenging task. This paper presents a systematic understanding of numerical modelling techniques for studying longwall coal mining with geological structures. The modelling techniques derived from conventional and advanced continuum and discontinuum methods were reviewed in detail with emphasiz on their mechanic's formulation and applications. This study suggests that the successful selection of a proper modelling technique should be based on the physical principle of longwall problem, texture and shape of materials, and mechanics formulation of the numerical program used for modelling. The paper's conclusions assist numerical modellers in quickly and properly selecting modelling technique for investigating a site-specific longwall problem.
... toppling and buckling) using the UDEC-GBM method [20,21]. Gao et al. improved the Voronoi logic and got the UDEC-Trigon block model of more realistic to model fracture patterns and simulated the process of crack development around roadways [22,23]. Zheng et al. studied the mechanisms of flexural toppling failure in anti-inclined rock slopes using UDEC-Trigon method [24]. ...
Slip and instability of coal-rock parting-coal structure (CRCS) subjected to excavation disturbance can easily induce coal-rock dynamic phenomena in deep coal mines. In this paper, the failure characteristics and influencing factors of CRCS slip and instability were investigated by theoretical analysis, numerical simulations, and field observations. The following main results are addressed: (1) the slip and instability of CRCS induced by excavation are due to stress release, and the damage of the rock parting is partitioned into three parts: shear failure zone, slipping zone, and splitting failure zone from inside to outside with slip; (2) the slip and instability process of CRCS is accompanied by initiation, expansion, and intersection of shear and tensile cracks. The development of the cracks is dominated by shear behaviour, while the tensile crack is the main factor affecting fracture and instability of CRCS; and (3) slip and instability of CRCS are characterized by stick-slip first and then stable slip, accompanied with high P-wave velocity and rockburst danger coefficient based on microseismic tomography.
... According to the SB01 drilling in the S1201-II working face, the intact rock mechanics parameters and rock quality designation (RQD) values of the coal measure strata were obtained. e elastic modulus, uniaxial compressive strength, and tensile strength of rock mass were obtained using equations reported in paper [17]. ...
Universal distinct element code (UDEC) is a simulation software based on the discrete element method, widely used in geotechnical mining. However, in the UDEC, when simulating large-scale excavation, the subsidence of the fractured zone is almost equal to the mining height, which makes the deformation value calculated in the study of gob-side entry retention too large. To solve this problem, in this paper, the double-yield constitutive model is applied to the whole caving zone to analyze the deformation and failure characteristics of surrounding rock along gob-side entry retaining by roof cutting. The results of the simulation are in good agreement with the result of drilling peeking (drilling observation by borehole televiewer) and field condition (observation and measurement in the field). Finally, by using this numerical method, the effects of roadway width, temporary support, and coal side support on the failure of the roof and the arc coal side are studied.
... Shen et al. (2008) revealed that prior to the roof falls, there are three processes emerge where firstly the seismic activity intensifies and then the stress reduces followed by roof displacement acceleration from their field testing. This phenomenon is demonstrated by DEM simulation (Gao et al. 2014b;Kang et al. 2018) and can also be found in our numerical results sequentially as cracks propagation, stress reduction and roof sag acceleration (see Fig. 23a, d). However, it is evident from Fig. 23d that there are some differences in these Fig. 23 Comparison of the fracture process and pattern of roof with weak and strong BASs, respectively. ...
The weak bedding sandwiched between rock layers is one of the crucial factors influencing the overburden failure with the extraction of coal resources in longwall mining. In this article, the physical model tests and discrete element modelling (DEM) were carried out to examine the effects of bedding adhesion strength (BAS) on the roof fracture with longwall coal mining. A novel model generation method in the universal distinct element code was proposed for the investigation of a propagating crack interaction with existing discontinuities. It is found that crack trajectory changes from deflection to penetration with the strength increase of a discontinuity. The stair-stepping fracture of overburden cantilever structure is more prominent for the rock layers with weaker BAS, which is in coincidence with the triangular block-based DEM result. Moreover, the mechanism of failure pattern influenced by BAS was clarified by the cantilever beam theory and fracture criterion of crack competition. With the outcomes above, the longwall coal mining-induced stratified roof fracture characteristics in engineering scale were revealed. Three quantitative regions were divided in terms of the strength ratio between bedding and rock layer. Finally, three aspects, the heights of the caved zone and fractured zone, the subsidence of ground surface, and the abutment pressure of roof affected by the BAS were discussed, respectively, with some suggestions recommended for the stability controlling of roof and safety production of coal mining.
... Sun et al. [3] proposed using a flexible bearing structure composed of retractable anchor and anchor cable as the main supporting mode based on the numerical simulation results, and then grouting reinforcement was adopted to control the deformation of soft rock roadway. However, the current research objects are mainly supported structures, which do not consider the self-bearing capacity of the surrounding rock, and the situation of mudstone softening on the floor cannot be reflected during the numerical analysis [4][5][6][7][8], which leads to the discrepancy between the simulation results and the real situation. is means that there is still room for optimization in the support scheme. ...
More and more attention has been paid to the supporting problem of deep soft rock roadway floor with long-term water immersion in recent years. However, the existing soft rock roadway support technology rarely takes into account the influence of the immersion softening phenomenon of the roadway floor and the self-supporting structure characteristics of the surrounding rock on the stability of the surrounding rock at the same time, and the influence of the creep characteristics of rock on the deformation zone of the surrounding rock requires further research on the nature and division of the self-supporting structure of the surrounding rock. In response to the issues mentioned, based on the loading and unloading properties of the surrounding rock of the soft rock roadway, a new concept of the internal and external self-bearing structure was proposed. The fact of water-immersed mudstone softening in the soft rock roadway floor was revealed through the field practice, and the shape of the internal and external bearing structure was determined based on the in situ monitoring results. Then, the instability mechanism of the internal and external self-bearing structure of the surrounding rock was analyzed, the position of the critical control point was calculated, and the key control technology based on the method of controlling floor heave by using double-row anchor cables to control the deformation of the roadway sides was put forward. Finally, the field industrial test showed that this support technology can effectively control the deformation and failure of soft rock roadway in the case of water immersion on the floor. This work can provide a technical reference for similar roadway support designs.
Discrete element method (DEM) has been widely used in studying fracture development of rock due to its ability to accurately depict particle interactions. In order to more intuitively describe the fracture characteristics of the main roof during coal seam mining using DEM, a novel numerical model for generating irregular particles roof (IPRM) in PFC3D is developed. In this novel model, irregular blocks are established using rblock, where the balls are placed to form irregular particles. Irregular particles use the flat-joint model, while the smooth-joint model is utilized between these irregular particles. The interlocking effect between the irregular particles of model can well restore the real failure characteristics of the main roof. Using the IPRM, five models of the main roof with varying thicknesses are created to investigate the failure characteristics with different thicknesses under uniform loads. The results show that the load-bearing capacity increases, and deflection decreases with the main roof thickness increasing. Additionally, the increase in main roof thickness leads to a shift in the failure pattern from “o-x” to “o- *,” accompanied by an increase in the fracture angle and the emergence of shear cracks. This change also leads to a transition of failure mode in the main roof from tensile failure to tensile-shear mixed failure. Finally, a mechanical model of the main roof is established, and the influence of different thicknesses and advance distance on the tensile stress and shear stress of the main roof is analyzed. It is found that the increase in the main roof thickness inhibits the development of tensile stress and promotes the development of shear stress, which is also the root cause of shear cracks and shear failure of the thick main roof. The study has theoretical guiding significance for ground control and is conducive to safe production of working face.
Traditional coal mining processes have increasingly raised concerns about roadways and surface damage. An innovative mining method, based on DSRC technology, has been proposed to address these issues. The DSRC technology involves using bilateral energy gathering devices to achieve roof cutting on both sides of the working face. The N00 method builds upon DSRC technology by incorporating additional auxiliary techniques to achieve roadways and surface protection during mining. With this method, mining N working faces can be accomplished with the advancement of 0 roadway and the retention of 0 coal pillar. The protection mechanism of the DSRC method on roadways and surface is analyzed through theoretical calculation, numerical simulation, and field engineering analysis. The N00 method with DSRC modifies the structure of the roadway roof by changing the long cantilever beam to a short cantilever beam. This modification reduces the stress on the roof and improves the stability of the roadway roof. The DSRC cuts off the stress transmission in a part of the roadway roof, reducing stress concentration on the coal side and enhancing its stability. The N00 method with DSRC increases the amount of strata bulking, utilizing bulking gangue to support the overlying strata, reducing the development height of the plastic zone in the gob, protecting the overlying aquifers, and mitigating damage to the surface ecological environment. Field monitoring results show that the N00 method with DSRC reduces the door-like support stress, NPR anchor cable stress, and deformation of the roadway, decreases stress concentration on the coal side, minimizes surface deformation, and reduces water inflow into the working face. Therefore, the purpose of protecting roadways and the surface is achieved.
Tensile and shear fractures are significant mechanisms for rock failure. Understanding the fractures that occur in rock can reveal rock failure mechanisms. Scanning electron microscopy (SEM) has been widely used to analyze tensile and shear fractures of rock on a mesoscopic scale. To quantify tensile and shear fractures, this study proposed an innovative method composed of SEM images and deep learning techniques to identify tensile and shear fractures in red sandstone. First, direct tensile and preset angle shear tests were performed for red sandstone to produce representative tensile and shear fracture surfaces, which were then observed by SEM. Second, these obtained SEM images were applied to develop deep learning models (AlexNet, VGG13, and SqueezeNet). Model evaluation showed that VGG13 was the best model, with a testing accuracy of 0.985. Third, the features of tensile and shear fractures of red sandstone learned by VGG13 were analyzed by the integrated gradient algorithm. VGG13 was then implemented to identify the distribution and proportion of tensile and shear fractures on the failure surfaces of rock fragments caused by uniaxial compression and Brazilian splitting tests. Results demonstrated the model feasibility and suggested that the proposed method can reveal rock failure mechanisms.
To propose a support stiffness based method for improving the stability of coal wall, static stiffness of the support and fracture development of the coal wall are monitored during retraction period of a panel in the second coal mine of Zhaogu coalfield. It is revealed that the range of support stiffness falls between 50 and 450 MN/m. In face length direction, support stiffness is characterized by three zones. It is small in the middle section while large magnitude is observed at two end areas. Fracture development of the coal wall is detected by geological radar, which reveal that fracture development is high in the middle section, and a decreasing trend is experienced from the middle to side areas. Such results imply face stability is strongly influenced by support stiffness. Based on such understanding, mechanical model for supporting system of the longwall face is established by considering the stiffness. The results demonstrate that roof pressure transferred onto the coal wall is negatively related to support stiffness. Then, in situ monitored relationship between support stiffness and fracture development of the coal wall is reasonably explained. Moreover, numerical modeling is carried out to validate the influence of support stiffness on face stability, and the results show the minimum support stiffness is proposed to be 100 MN/m for the target panel. The study provides a new method for selecting support stiffness in longwall panel with similar geological conditions.
Grouting is an effective method to improve the integrity and stability of fractured rocks that surround deep roadways. After years of research and practice, various theories and a complete set of grouting technologies for deep roadways with fractured rocks have been developed and are widely applied in Chinese coal mining production. This paper systematically summarizes and analyzes the research results concerning the theory, design, materials, processes, and equipment for the grouting and reinforcement of fractured rocks surrounding deep roadways. Specifically, in terms of grouting methods, pregrouting, grouting‐while‐excavation, and postgrouting methods are explored; in terms of grouting theory, backfill grouting, compaction grouting, infiltration grouting, and fracture grouting theories are studied. In addition, this paper also studies grouting borehole arrangement, water‐cement ratio, grouting pressure, grouting volume, grout diffusion radius, and other grouting parameters and their determination methods. On this basis, this paper explores the physical and mechanical properties of organic and organic‐inorganic composite grouting materials, and assess grouting reinforcement quality testing methods and instruments. Taken as the field cases, the application of pregrouting in front of heading faces, grouting‐while‐excavation, and postgrouting in the Kouzidong coal mine are then introduced, and the effects of the grouting reinforcements are evaluated. This paper proposes a development direction for grouting technology based on problems existing in the grouting reinforcement of fractured rocks surrounding deep roadways.
As a kind of newly emerged excavation equipment, the TBMs play an increasing important role in underground coal mines due to their high penetration rates, while the failure and large deformation of large cross-section TBM assembly chambers need to be controlled under combined action of high ground stresses and live loads generated by heavy TBM parts. In this research, the mechanical parameters of rock specimens and rock quality designation (RQD) were used to estimate mechanical parameters of rock masses. The numerical tests were conducted and the micro-parameters of surrounding rocks in UDEC were calibrated. A numerical model of the TBM assembly chamber was established and the disturbance characters under the original support scheme after excavation were studied. Moreover, the influences of live loads generated during the hoisting and assembling operations of TBM parts were taken into account. An optimized support scheme was proposed and examined by numerical simulation. The in situ monitoring works were conducted to verify the simulation results. The monitoring results are in good agreements with the simulation results and the optimized support scheme satisfied the demands of stability controlling of the TBM assembly chamber.
In view of the problem that the broken stope roof under repeated mining is prone to end face roof leaks, this paper’s research background is the close distance coal seams mining of a mine in Guizhou province, China. The impact of stope roof structure movement on the immediate roof was elucidated from the analysis of stope roof structure movement under repeated mining employing a combination method of field measurement, physical simulation, theoretical analysis, and numerical simulation. As a result, the mechanism of end face roof leaks under repeated mining is investigated, and the corresponding control measures are proposed. The results show that the cantilever beam structure instability of the stope roof under repeated mining is an important reason for the end face roof leaks and support crushing accident. The fracture forms of cantilever beam structure are divided into direct fracture and hinged structure fracture. The direct fracture is affected by the support working resistance. The cantilever beam hinged structure is divided into sliding and rotational instability. The periodic weighting interval is small and the fracture degree is large under repeated mining, which leads to the sliding instability. It is concluded that reasonable support working resistance is the important reason to keep the cantilever beam structure stability. Therefore, the stope roof leakage control countermeasures under repeated mining are proposed. The UDEC numerical simulation is used to verify the above analysis again, so as to provide the basis for the end face roof leaks prevention and control under repeated mining.
The roadway direction is often parallel to the direction of the maximum principal stress for the structure arrangement of underground roadway. To study the instability mechanism of roadway under the maximum principal stress, a series of uniaxial compression acoustic emission (AE) experiments were carried out on hollow red sandstone and granite rock samples with different bore diameters. The test results suggest that the oblique shear or X-shaped shear failure occurs in hollow rock samples, and the failure process is accompanied by a sudden jump in accumulative energy of AE signals, at the same time, the ringing count arrived at peak value. The increase in bore diameter haves contributed to the decline of ringing count and accumulative energy. A statistical damage constitutive model of Weibull distribution was established for hollow rock samples. On the basis of optimizing the damage variable, a damage constitutive model based on AE ringing count was obtained, which can better describe the damage state and stress state of rock samples. The AE accumulative energy is selected as the state variable, and the failure criterion model was finished based on the cusp catastrophe theory to estimate the rock sample failure. The judgment results show that the failure moment of rock sample is accurately included in the judgment time interval. This research can provide reference and technical support for the stability control and monitoring of underground engineering roadways.
The double-roadway layout system, which is extensively applied in large mines, has the potential to significantly balance excavation-mining and improve mine ventilation and transportation capacity. However, the coal pillar in the double-roadway layout system is easily destabilized due to the disturbance of repeated mining, which has a significant impact on the safety and reliability of coal mines. This paper takes the coal pillar and its supporting structure of the double-roadway layout system as the research object, establishes a UDEC trigon numerical calculation model, and systematically corrects the input parameters, while explaining the excavation method of roadways and the simulation method of the supporting structure element. The numerical simulation results show that under the conventional support intensity conditions, the internal damage of the coal pillar during the excavation period is about 20%, while the internal damage to the coal pillar develops to 55% throughout the first-panel mining. During the disturbance of repeated mining, the damage in the coal pillar increased to 90%, and the coal pillar was already in a state of failure. Under the combined control of rock bolts and counter-pulled anchor cables, the coal pillar damage does not change significantly during the excavation and first-panel mining. During the disturbance of repeated mining, the damage of the coal pillar is reduced to 63%. There is a certain low damage area in the coal pillar, which can ensure the stability of the coal pillar and its supporting structure as a whole. Furthermore, the on-site monitoring results show that the maximum value of the floor-to-roof and rib-to-rib convergence of a W1310 tailgate during the repeated mining disturbance stage is 730 and 620 mm, respectively. The findings of this study give an approach to—as well as estimated values for the design of, including its “small structure” control technical parameters—the double-roadway layout system.
The large deformation issue in soft rock mass is a worldwide difficulty that has puzzled tunnel engineering for a century and a half. Previous studies have mainly focused on a large deformation mechanism, prediction, and control of soft rock with single lithology, while there are limited studies on the tunneling-induced large deformation in monoclinic and soft-hard interbedded rock strata. This paper studies the deformation law, evolution process, and failure mechanism of a tunnel excavated in monoclinic and soft-hard interbedded rock mass by onsite measurements in the No. 3 inclined shaft of the Muzhailing Tunnel, which is a key control project of the Lanzhou-Haikou National Expressway. The achieved results indicated that the tunnel deformation was characterized by significant asymmetry, the maximum deformation in the cross section was observed to occur in the normal direction of the bedding plane, and the side of the larger deformation was consistent with the dip direction of the strata. The vertical displacements were greater than the horizontal displacements except for the right foot of the middle bench, and the primary lining at the upper left of the tunnel had intruded into the space of the secondary lining and had to be demolished and reconstructed. The displacement speed and amount mostly reached the peak in the construction stage of the middle bench, and its displacement amount accounted for about half of the total displacement. The spatiotemporal curves of large deformation could be divided into four stages according to the excavation time and the distance from the tunnel face. Two types of exponential functions were used to fit and predict the tunnel displacement, which displayed good applicability. The disaster evolution process of large deformations was summarized into five stages on a macro scale: premise, gestation, development (including four substages), occurrence, and treatment of large deformation. The failure mechanism of monoclinic and soft-hard interbedded rock mass is mainly the bending-tensile failure of thin-layered strata.
Simplification of complex geologic systems has been a necessary hallmark of geological engineering research and design to date. However, oversimplification and subsequent over-application of existing methods leaves room for significant improvement in our understanding of rockmass response to excavation. Although indisputable advancements have been made in increasing the safety of underground workings, falls of ground continue to injure or kill personnel and delay production. The overapplication of existing simplified methods is particularly problematic in discontinuous and laminated systems, where the response to excavation can be anisotropic and significantly impacted by the orientation, intensity, and condition of discontinuities. With the advancement of computational power and numerical modeling techniques, more of the mechanical complexities associated with discontinuous systems can be explicitly considered. Therefore, the goal of this research is to identify the geomechanical considerations for a wide range of discontinuous and laminated geologic conditions that should be incorporated into analytical and empirical methods to increase the safety and productivity of mining and civil works. This thesis focuses on addressing and overcoming two of the most significant simplifications often employed in the design of flat-roof excavations: assuming that the overburden has no self-supporting capacity, and representing discontinuous systems as continua. To that end, this research utilizes the explicit discrete element method (DEM) to identify and account for the relevant geologic and mining conditions that control local and global stability. Model complexity and scale is increased incrementally, and model results are compared to existing, well-established analytical and empirical methods to validate, confirm, or frame the implications of the numerical results and their relationship with “reality”. The first objective of this thesis is to evaluate roof self-stability and stress arching capacity through application and enhancement of the voussoir beam analog. Gaps in existing analytical calculations are identified and addressed through the methodical variation of geometry, material properties, and boundary conditions in explicit DEM voussoir beam numerical models. An adjusted voussoir beam analog is developed that can account for novel aspects of complexity such as post-peak material behavior, horizontal stress, and layered roofs that are passively bolted. The adjusted voussoir beam analytical method is then applied to a case study of the Bondi Pumping Chamber excavation in Sydney, New South Wales, Australia. The second objective of this thesis is to analyze roof self-supporting capacity and bolted stability through a parametric sensitivity analysis of 8,640 unique explicit DEM models of hypothetical coal-mine entries conducted with a particular focus on discontinuity properties. Additional considerations include in-situ stress magnitude and horizontal stress ratio, as well as material stiffness, strength, and anisotropy. Model inputs are utilized to assign a Coal Mine Roof Rating (CMRR) value to each model case, and the Analysis of Roof Bolt Systems (ARBS) is subsequently used to assess the reliability of the model results and focus future statistical analysis. Multivariate binary logistic regression is used to identify the statistically significant parameter inputs that determine the probability of a stable roof condition in unsupported and bolted models. Recommendations such as adjusting the cohesion-roughness rating and consideration of joint orientation in CMRR, as well as accounting for in-situ horizontal stress ratio in ARBS, are posited. The last objective of this thesis is to identify how excavation roofs and pillars are mechanically linked. A calibrated, confinement-dependent coal pillar constitutive model is combined with the significant controls on roof stability identified through the course of this study to assess pillar-overburden interaction in single-entry and multi-entry models. Entry-scale models are used to identify the interaction between roof stress arching capacity and pillar confinement, and panel-scale models are subsequently developed to incorporate in-situ complexities such as panel width-to-height ratio, lithologic heterogeneity, and depillaring to assess overburden stress arching capacity and pillar response. Lastly, the panel-scale model results are compared to state-of-practice analytical and empirical methods such as tributary area theory (TAT), the Analysis of Retreat Mining Pillar Stability (ARMPS), the abutment angle concept, and the Mark-Bieniawski pillar strength equation. Results confirm that properties that increase stress arching in the overburden tend to decrease pillar loads and increase pillar strength. The results of this study identify that increasing both the accuracy and applicability of existing analytical and empirical methods, as well as our holistic understanding of flat-roof excavation stability requires mechanically coupling the pillars to the roof and floor. Without this explicit consideration, state-of-practice and state-of-knowledge cannot advance towards both safer and more efficient excavations.
Bending failure is a common failure mode of layered rock mass. Making clear the mechanical behaviors and energy evolution characteristics of layered rock mass, it is beneficial to prevent geological disasters caused by the bending deformation of layered rock mass. In this study, the mechanical behaviors and energy evolution characteristics of hard and soft rocks by conducting the three-point-bending (TPB) test with different loading rates were investigated. The results show that as the loading rate increases, both the peak load of hard and soft rock increases, the peak displacement of hard rock decreases, while the peak displacement of soft rock increases. The horizontal crack width at the bottom of the sample of hard rock is greater than that of soft rock, but the instantaneous crack widths show opposite results. Both the failure pattern of hard rock and soft rock are the typical tensile fracture, yet the fracture surface of hard rock is denser and smoother than that of soft rock. For hard rock, the total input energy, elastic energy, and dissipated energy increase with the increase of loading rate. For soft rock, however, the total input energy and elastic energy increases, while the dissipated energy decreases. Under the TPB test, the peak load, displacement, instantaneous crack width, total input energy, elastic energy, and dissipated energy of both hard rock and soft rock present linear relationships with the common logarithm of the loading rate.
Understanding coal rib geomechanics is essential for improving rib stability and eliminating fatality and injury trends due to rib failures. There are presently no standardized rib control practices available in most countries. In light of this observed dearth, this investigation aims to improve understanding of rib failure mechanisms using the distinct element modeling (DEM) technique. DEM is chosen because of its superior advantage to explicitly represent discontinuities and their constitutive behaviors, besides that of the intact rock matrix. To analyze the rib stability, a numerical monitoring protocol is implemented to monitor the deformation characteristics of the coal rock mass as its strength is gradually reduced and the deformation and safety factors are established. A number of scenarios were considered in the modeling process, including a non-cleated rib, a cleated rib, and the interaction between support and coal mass. The main conclusions drawn from the study were that the rib failure process initiated with tensile and shear cracks which coalesced to form predominantly sub-parallel tensile fractures to the rib line; and joints and defects in the rib limit fracture development and propagation. The depth of fracture was found to be ~ 1.14 m for the cleated rib and ~ 1.40 m for the non-cleated rib. The depth of softening (DOS) for the cleated and non-cleated ribs was ~ 1.80 m and ~ 1.60 m, respectively. Also, the results demonstrated the capability of DEM-based bonded block models (BBM) in explicitly capturing the rock and support interactions which makes this solution to be suitable for investigating coal rib stability and support requirements.
There is an increasing number of TBMs applied in coal mine roadway excavation projects due to the higher safety, faster tunnelling speed and fewer labor works. Nevertheless, the experiences gained from TBM tunnelling projects of transportation or water conveyance tunnels are not well applicable in underground coal mines because of their special engineering and geological conditions. This paper proposed a case study of the excavation disturbance behaviors and support design of TBM-excavated roadway in Zhangji coal mine in Anhui Province, China. Mechanical properties of roadway surrounding rocks were obtained based on parameters of intact rocks and Rock Quality Designation (RQD) values of rock masses. The Universal Discrete Element Code (UDEC) software was used to establish numerical model of the roadway and micro-parameters are also calibrated. The excavation disturbance behaviors of unsupported roadway were firstly simulated. Moreover, the excavation disturbance behaviors and rockbolts loading characters under two different support schemes were studied and compared based on simulation and monitoring results. In this case, simulation results are in good agreement with monitoring results. Research results and experiences obtained from the working site suggest that installing rockbolts perpendicularly to rock surface can better satisfy the requirements of TBM tunnelling in weak grounds of coal mine roadways.
The roadways with top coal are a common type of roadways in thick coal seam mining. After excavation of such roadways, the top-coal rock mass is often loose and broken, and the roof deformation and failure are serious. Thus, the stability control of such roadways is a major concern of engineering design and construction personnel. In this paper, we take the typical roadways with top coal in Zhaolou coal mine of China’s Juye mining area as an example and propose a composite failure mechanism for roadway roof supported by rock bolting. The roof stratification and nonlinear failure characteristics of rock masses are incorporated in it. Then, a simplified design method of roof bolting support is proposed on the basis of the upper bound theorem. Furthermore, this paper shows the influence laws of varying parameters on the required supporting force of roof bolts and provides several practical recommendations for engineering applications. In addition, the effectiveness of the proposed method in this paper is further verified through a specific case analysis on a site typical roadway. The research work in this paper can provide theoretical guidance for support design and construction in roadways with top coal.
Rock reinforcement design necessitates a clear understanding of the initiation and propagation of cracks in the bolting system. To better understand the mechanism, we performed laboratory pull-out tests on resin-encapsulated rock bolts. Numerical models on a 1:1 scale were built using the discrete element method. The microscopic parameters of the model were calibrated based on unconfined compression tests and ring shear tests. The model allows us to visualize the progressive failure of a bolting system strengthened with resin-encapsulated rock bolts and elucidates the role of the anchorage length in controlling crack propagation. Comparisons indicate that the physical and numerical results are consistent. The results show that increasing the anchorage length improves the bondability and strength of the bolting system and restrains the complete debonding of the rock bolt from the cement mortar. At the same time, it also facilitates the conversion of shear cracks to tensile cracks. However, the shear crack of the resin is the dominant effect in the bolting system. In addition, the results reveal how factors such as boundary confining pressure, resin thickness, and cement-mortar strength affect crack initiation and propagation.
NIOSH researchers have identified a pattern of fracture zone development that suggests an explanation for fracture formation around rectangular openings in underground mines. This pattern is characterized by shearing and dilation that result in either faults or in the repeated formation and propagation of en echelon fractures from sites of tension. Two computer modeling codes, Fast Lagrangian Analysis of Continua (FLAC) and Particle Flow Code (PFC), were used to model different aspects of this pattern. Use of very small elements with FLAC enabled the identification of sites of initial tension near the skin of square-cornered rectangular openings, while PFC allowed the initiation and progressive development of fractures from these sites to be followed as the fractures evolved. Such studies can lead to a greater understanding of how roof support can be better selected and installed for specific conditions in underground mines prone to roof falls and rock bursts. These studies may also lead to modifications of corner and opening shapes that could be incorporated into mine designs to produce more stable mine openings and reduce the risks of rock falls and rock bursts.1. INTRODUCTIONMining-induced fractures are typically either hidden from view or become obscured when fractured rock around mine openings collapses or is ejected in a rock burst. Consequently, knowledge of the distribution, geometry, and extent of mining-induced fractures has been limited. However, fracture patterns that suggest mechanisms of fracture formation are occasionally seen where new crosscuts intersect old ones. The old fractures are exposed in the walls of the new opening, but well-described examples are rare in the literature. Most investigators agree that extension fractures form parallel to the direction of maximum principal stress. Recently favored explanations for how decimeter and longer (macroscale) fractures develop have mainly involved the proliferation, interaction, and coalescence of smaller fractures. Fairhurst and Cook [1] proposed that stress-induced microcracks first form an incipient cleavage parallel to the surface of a mine opening. These microcracks then grow into larger fractures as a result of the applied stress. These fractures shorten with depth, and their ends define a shape concave toward the opening that represents the limit of breakage in the case of a rock burst, roof collapse, or pillar hour-glassing. However, in a field example where thin layers were displayed at the margin of a rock burst breakout, White [2] concluded that the closely spaced fractures had not extended across the entire volume of the ejected rock, but were present only near the periphery of the breakout. Other examples that support this scenario are described in White et al.[3] Stacey [4] proposed that failure about mine openings will occur if extension strain reached a certain value, which he considered a material property. He noted that extension strain is highest near the corners of rectangular openings and suggested that failure begins at these locations. He proposed that the extent of failure is determined by how much rock around the opening exceeds the requisite extension strain. For rectangular openings, critical extension strain is deepest at the midpoint between corners and its limit duplicates the concave shape identified by Fairhurst and Cook [1] and commonly seen after roof falls and rock bursts. However, Stacey did not differentiate extension resulting from tension from Poissonresponse extension.
The National Institute for Occupational Safety and Health (NIOSH) evaluated microseismic activity from three field sites to compare and contrast the characteristics of microseismic emissions from very different geologic, stress, and mining environments. Recently, NIOSH has embarked on a research program to evaluate the use of microseismic monitoring information to identify roof fall failure processes and to assess its potential to warn of unstable roof conditions. Large roof instabilities, such as roof falls and certain roof caving events, have proven difficult to anticipate representing an increased risk to miners working in these inherently hazardous areas. When local failure processes are better understood, appropriate control measures can be engineered to mitigate these hazards. This study used microseismic emissions to help identify three local rock failure processes. It was also shown that analysis of microseismic emissions can aid in assessing the degree of instability associated with these local rock failure processes.
Techniques used to safely control the roof in UK coal mines for some 10 years have been applied to a South African coal mine. This new approach known in the UK as advanced rockbolting technology, is based on applying four fundamental principles: ▶ Understanding the roof failure mechanisms ▶ Using an effective roof support system ▶ Designing this support using measurement ▶ Monitoring the performance of the system. The paper summarizes the approach in detail and describes how it is being applied by Anglo Coal, giving results obtained to date. Investigations at a number of South African coal mines, including stress measurements at two Anglo Coal mines, have confirmed that the mode of roof failure (lateral shearing due to horizontal stress) is the same as in other coalfields worldwide, and aspects of current world best practice, including the use of advanced technology rockbolting, are therefore relevant in South Africa. The stress measurements indicated a high level of stress field anisotropy and further investigation of stress conditions in South African coal mines is recommended. The most effective bolting system to resist shear failure is one with high bond strength and stiffness. Short encapsulation pull testing of existing South African systems confirmed that they have low bond strength and stiffness. An improved rockbolt system with the required performance, and features allowing rapid installation and installation quality and performance audit, has been developed and is currently under full-scale trial. Design by measurement and routine monitoring procedures including the use of a rotary telltale device are also under trial. It is anticipated that South African coal mines will be able to obtain significant safety and productivity benefits from the application of this technology.
The National Institute for Occupational Safety and Health (NIOSH) has continued the research role of the former US Bureau of Mines (USBM) to engineer techniques that will reduce the hazards in the mining work place associated with coal bumps. Recent research focused on a longwall coal mine in Utah with overburden greater than 750 m (2,500 ft) containing several massive sandstone units. The primary field instrumentation at the site was three-dimensional, full waveform, autonomous microseismic arrays placed underground and on the surface in order to surround the active multi-panel longwall district. The purpose of these arrays was to help investigate the strata mechanics associated with the redistribution of stress and the associated gob formation of the longwall. Specifically, the s eismic ar rays were used to determine the timing and location of the failure in t he strata surrounding the active mining. Overall 13,000 seismic events were detected and located with on-site processing during the five months the panel was being mined, including a magnitude (ML) 4.2 event. Of these, a smaller subset of 5,000 well-located events was selected during post-processing to form a consistent data set for analysis in this paper. From this data set, it was observed that the seismic events generally occurred in advance of the longwall face, both above and below the panel, consistent with failure of the strata in the forward stress abutment zone. Also, the occurrence of the ML 4.2 seismic event within 150-180 m (500-600 ft) of a deep cover longwall face with no associated bump caused a re-evaluation of the nature of the connection between seismic activity and coal bumps.
Roof stability in gateroads is a long-standing issue in many of the underground mines in Australia that use longwall extraction methods, due primarily to a significant increase of vertical stresses ahead of the longwall face. Although numerous studies have been done in the past, the process of roof rock deformation and breakage prior to and during a roof failure in an actual mining environment is still being debated. This paper describes a new integrated roof monitoring system and the results from applying this system in an Australian underground coal mine. The system integrates displacement, stress and seismic monitoring. It has been applied to two roadways in an Australian underground coal mine during two field experiments. The key roof behaviour identified by the integrated monitoring package during the two field monitoring experiments is reported and discussed in this paper.The experiments were conducted in the “tailgate” roadways that are adjacent to the caved zone, or “goaf”, of the previously mined panels. It was found in the experiments that, prior to roof falls, roof displacement accelerates whereas the horizontal stresses reduce. Seismic activity intensifies before major roof displacement or stress changes are evident, and subsides in the later stage of roof failure when large roof displacement is visible. The seismic resonance frequencies decrease during roof failure development. The field monitoring studies have also identified a number of quantitative and site-specific roof fall precursors potentially useful for roof fall prediction and prevention.
A deep longwall coal mine was instrumented with a three- dimensional microseismic system in order to help determine the exact strata mechanics associated with the rock failure, redistribution of stress and the associated gob formation from the longwall. Overall, 5,000 well calibrated seismic events were recorded during the mining of the panel. Analysis of these events showed a close correlation between the seismic activity and advance rate, and that the majority of the recorded seismic activity occurs in the immediate area of the advancing longwall face.
The National Institute for Occupational Safety and Health (NIOSH) has continued the research role of the former U.S. Bureau of Mines to develop techniques that will reduce the hazards in the mining work place associated with coal bumps. Current research focuses on both analyzing historical seismic data from bump-prone operations and utilizing a mine-wide seismic network to investigate the exact strata failure mechanics associated with bump-prone geology. The anticipated outcome of this research will be reduced bump incidences through advanced engineering concepts and designs which implement the new understanding of strata behavior. The analysis of the historic seismic data consists of correlating observed mining seismicity with the geologic and geometric parameters at the sites. The primary seismic parameters are the timing, location and magnitude of a recorded seismic event. These parameters are correlated with such mining parameters as: the overburden, the size of the immediate gob, the size of the district gob area, etc. This detailed analysis of historical seismic data has provided an informative quantifiable relationship between many of the specific mining parameters and the induced seismicity. The second aspect of the coal bump research is the instrumentation of an appropriate field site to determine the main roof, floor, and gob behavior associated with bump behavior. The chosen field site is a deep-cover longwall mine with competent geology in a historically bump-prone area. The primary field instrumentation is a three-dimensional, full- waveform, seismic array with both surface and underground sensors surrounding an active multi-panel district. The purpose of this seismic array is to determine the timing, the exact location, and the mechanism (tensile fracture, bedding plane slip, etc.) of the failure of the strata surrounding the active and multi-panel gobs. The preliminary results presented in this paper help to define the strata failure areas around the longwall panel.
For the first time in an underground stone mine, the relationship between roof movement and microseismic emissions was examined in conjunction with two distinct roof fall areas. As roof monitoring increases in acceptance and monitoring technology advances, the goal of providing reliable roof fall detection systems to enhance the safety of underground mine workers moves closer to reality. Instrumental to reaching that goal is the ability to interpret accurately and completely roof movement and microseismic emissions, which can serve as precursors to roof falls. This paper examines the capabilities of convergence and microseismic monitoring systems to better understand roof rock failure mechanics and to anticipate roof falls. An understanding of how these techniques are used and how they interact with each other is critical in developing the most effective ground control strategy for our nation's mines.
One of the critical design problems involved in deep tunnelling in brittle rock with continuous excavation techniques, such as those utilizing tunnel boring machines or raise-bore equipment, is the creation of surface spall damage and breakouts. The mechanisms involved in this process are described in this paper. The onset and depth of damage associated with this phenomenon can be predicted, as a worst case estimate, using a factored in situ strength value based on the standard uniaxial compressive strength (UCS), of intact test samples. The factor applied to the UCS to obtain the lower bound in situ strength has been shown repeatedly to be in the range of 0.35–0.45 for granitic rocks. This factor varies, however, across different rock classes and must be determined or estimated for each class. Empirical guidance is given for estimating the in situ strength factor based on the UCS for different rock types and for different descriptive parameters. Laboratory testing procedures are outlined for determining both this lower bound strength factor and the upper bound in situ strength. This latter threshold is based on the definition of yield based on crack interaction. These techniques are based, in part, on theoretical principles derived from discrete element micromechanical experimentation and laboratory test results. The mechanisms that lead to in situ strength drop, from the upper bound defined by crack interaction and the lower bound limited by crack initiation, are described. These factors include the influence of tunnel-induced stress rotation on crack propagation, interaction and ultimately coalescence and failure. A case study illustrating the profound impact of near-face stress rotation is presented.
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