Movement and failure law of slope and overlying strata during underground mining

Transcription

1 Journal of Geophysics and Engineering J. Geophys. Eng. 5 (08) (3pp) ovement and failure law of slope and overlying strata during underground mining Lv Xiangfeng,, Zhou Hongyuan, Wang Zhenwei 3 and Cai Yue 4 Beijing Key Laboratory of Underground Construction Forecasting and Warning, Beijing 00037, People s Republic of China Geotechnical Engineering Research Center, Institute of unicipal Engineering, Beijing 00037, People s Republic of China 3 ine Safety Technology Branch of China Coal Research Institute, Beijing 0003, People s Republic of China 4 School of Civil Engineering, Beijing Jiaotong University, Beijing 00044, People s Republic of China szgcyjylvxiangfeng@63.com Received 7 November 07, revised 8 arch 08 Accepted for publication arch 08 Published 9 ay 08 Abstract Overburden deformation has a considerable influence on slope stability during underground mining works. A simulation experiment for underground mining is performed in this study. The movement of the overlying strata caused by mining during the mining process is observed using a three-dimensional optical displacement monitoring system. Based on the simulation experiment results, a particle contact-based meshfree method (PC) is proposed. The elastoplastic analysis and sliding process simulation of the slope are realized by introducing the softening ohr Coulomb model and maximal tensile stress model into continuous media elements. The deformation process and slope failure are analyzed by using the PC numerical simulation method. The results showed that the overburden gradually degraded as the work face continuously pushed forward. Furthermore, the basic roof bore the force generated by the upper rock mass. Hence, a rock beam force distribution developed. The farther the distance from the coal seam, the smaller the subsidence. The rock strata subsided and induced a beam structure effect when the work face continued to advance past the design stop line, causing a rock mass partial settlement on the slope and a serious horizontal movement on the free face. This behavior significantly affected the slope stability, even inducing cracks and potential slip and sliding surfaces. The numerical simulation results indicated that the deformation process of the overlying strata is as follows: the roof immediately starts to collapse the rock mass of the main roof collapses overlying strata generate an abscission layer a fracture of the free surface develops the overlying strata begins to sink. eanwhile, the slope deformation and failure process is as follows: overlying rock mass exhibits shear failure the overburden is completely sheared the overburden collapses the slope surface exhibits a tensile crack the stepped crack is connected the slope begins to slide to the outside. The research results provide guidance for underground mining. Keywords: underground geotechnical engineering, three-dimensional optical photogrammetry system, displacement field, PC numerical simulation method, slope stability (Some figures may appear in colour only in the online journal). Introduction The main difficulty encountered in underground mining relating to underground geotechnical engineering is slope instability caused by excavation (Wang et al 998, Zhu 99). However, underground geotechnical mining always up-scales to 00 m or higher. Therefore, the scope of such work is very large, and obtaining a unified solution to 74-3/8/ $ Sinopec Geophysical Research Institute Printed in the UK

2 this problem in practice is difficult (Song et al 00). This study used a three-dimensional (3D) optical displacement monitoring system and a particle contact-based meshfree method (PC) numerical simulation method to examine the law governing the overlying strata deformation during the work face advancing process. The deformation process of the overlying strata and the slope during the work face advancing process was also monitored. The deformation process and the slope failure are analyzed after a simulation experiment by using the PC numerical simulation method. Finally, a theoretical basis for the mining project is provided for use in practical applications.. Similarity condition and similarity model.. Similarity condition... Geometric similarity. According to the practical situation and test conditions, the similarity ratio between the similarity model and prototype ratio is /50 (Huang et al 998, Chen et al 008), and the formula is as follows: a H LH = T. ( ) In the formula, L H and L represent the prototype and model length (m), respectively. α L is the geometric similarity constant.... otion similarity. The requirements of the model and prototype at all corresponding points of movement are similar, and require that the corresponding points of speed, acceleration, and movement time, are of a certain proportion. The relation between the time similarity constant α t and the geometric similarity constant α L can be obtained from the general equation of Newton s second law, F = ma. First, acceleration a is represented by dimensions [T] and [L], as shown below: a a H H LH L = a = ( ) L t H a a H H LH t =. ( 3) L t H To study the failure and caving of the rock mass under the action of gravity and inertia, we use the following equations: a = a = g ( 4) so: and: H LH t 5 L t = ( ) t t H H LH = ( 6) L obtained: th at = = a L = 5.8. ( 7) t In the formula, α t is the time ratio, t H is the time required for prototype movement, and t is the time required for similarity model movement. The actual working times of the spot are 8 h, while the actual working hours of the model are calculated according to the formula: 8 t = = 0.5 hours = 30 minutes. 5.8 According to specific laboratory conditions, the working face is simulated. This simulation is carried out at a speed of m per minute, until the stop line...3. Stress similarity. According to similarity theory, the stress similarity constant is: as = aga L =.8 50 = 450. In the formula, K is the stress similarity constant, and k is the bulk density constant, α γ =.8... Similarity material In the experiment, the apparent density and compressive strength were chosen as the necessary similarity parameters of the prototype and the model. Deformation, shear strength, elastic modulus, and Poisson s ratio are the unnecessary similarity parameters. Silica sand, calcium carbonate, gypsum, and other materials in accordance with the specified compressive strength, were selected as the aggregate. Cement, lime, kaolin, paraffin, and distilled water with a retarder were selected as the cement. Similar materials were mixed in proportion to replace the prototype material. For this study, a similar material apparent density of.5 g cm 3, a silica sand grain size of mesh, and a mica powder isolation layer were selected. The rock mass material parameters are shown in table, and the ratio of the similarity material is shown in table..3. Similarity model construction and excavation.3.. Similarity model construction. Based on the geometric similarity constant and the prototype model size, the thickness of each layer of a similar model was calculated (as tables and show). Then, the mixed similarity material was sequentially filled in the channel steel frame (each time filled 5 mm 0 mm.). In the experiment, the physical model platform size was 5000 mm 300 mm 000 mm, with a removable iron channel 00 mm wide used to fixed the front and back of the model. The physical model is shown in 639

3 640 aterial layer (from top to bottom) Component Thickness/m Table. Some mechanical parameters of coal and rock. Compressive strength/pa Internal friction angle/ Cohesion Elastic modulus/pa /Pa Bulk density/kg m 3 Poisson s ratio Backfill Loess Weathered Sandstone Clay mineral # coal Sandy mudstone 8 Fine Silty Gray # coal udstone # coal iddle-fine J. Geophys. Eng. 5 (08) 638 L Xiangfeng et al

4 Table. aterial proportion table. aterial layer Component Thickness/m Proportion number Sand binder ratio Cementing material oisture Compressive strength/pa Lime Gypsum Prototype odel Backfill / / Loess / / Weathered / / Sandstone / / Clay mineral / / # coal / / Sandy / / mudstone 8 Fine / / Silty / / Gray / / # coal / / udstone / / # coal / / iddle-fine / / figure (C). The specific steps to establish the model are as follows: () Install the template. Close the side of the model completely with the template, and then mark the layered position and the position of the slope on the template according to the model size. ake a line marking to use as the dividing line of the filler material. On the other side of the model, build the model while installing the template. () Preparation of similar materials. Weigh the sand, lime, gypsum and water according to the mix proportion. ix the sand, lime and gypsum evenly, and then add clear water. After uniform stirring, stand for h under standard curing conditions. (3) Fill the model. Fill the similar materials according to the sequence of the rock strata, filling 5 0 mm each time, with a compaction of 00 0 times. Using the line marking as a reference line, the relative error between the filling rock strata thickness and the calculated rock strata thickness is less than or equal to mm. After the single rock strata is completed, mm thick mica powder is sprinkled over it. Then, continue to fill the other rock strata in the same way until completed. (4) Demould. aterials reach their stable strength after seven days, at which time the template installed before and after the model can be removed..3.. Similarity model excavation. In the experiment, an underground mining method is adopted, mainly mining the no.4 and no.9 coal seams. During mining, the coal seams are reserved for coal pillars of 0.5 m. The work face advances toward the slope, and the mining rate is m per minute, where coal seam no.4 has a total mining length of.048 m and is mined twice. The mining length is.55 m and m, respectively. Coal Seam no.9 has a total mining length of.048 m, and was mined a total of three times 0.48 m for the first,. m for the second and 0.68 m for the third. 3. 3D optical photography method 3.. easurement equipment The traditional method of monitoring the movement and deformation of rock formations and the Earth s surface involves a physical or mechanical measurement. However, this approach has many shortcomings, such as the cumbersome installation of the observation device or sensor, heavy workload, and limited number of sampling points (Behrooz et al 05). The XJTUDP 3D optical photogrammetry system (figure ) is an industrial, non-contact, optical, 3D coordinate measurement system, which is also known as the digital industrial close-range photogrammetry system. This system can accurately obtain discrete 3D coordinates of a target point and is a portable, mobile, 3D, optical measuring system (Song et al 00). The single-point bit accuracy is mm, whereas the accuracy is /70000 / The actual observation error of 0 mm corresponds to an error of 0.04 mm in the model for a similar material model with a model ratio of :50. Due to its obvious advantages, this 64

5 method has been applied in many disciplines, such as solid mechanics, biomechanics, engineering mechanics, textile mechanics and industrial testing. However, this method also has disadvantages. It is noted that measured object surface non-uniform illumination or light intensity changes will result in measurement error (Zi et al 05, Sun 07). In order to avoid this problem, an incandescent lamp was used as the light source in this study. 3.. easuring point arrangement In accordance with previous studies, and the problems to be solved, the experimental grid line is laid conforming to a size of 00 mm 00 mm, and the displacement monitoring points are set at the intersection of the vertical and horizontal grid lines (Zhu et al 00, Ding et al 04, Cheng et al 07, Liu 07). The displacement is monitored by an XJTUDP 3D optical photogrammetry system. The system coordinates of the coded and non-coded points are also obtained (Jing et al 007, Jin and Ai Yan 009). Note that each coding point is a global control solution point. A sample coding point is indicated by the larger black marker shown in figure. In order to monitor the distribution and variation of the support system stress during the process of the working face, the BW-5 pressure box is pre-embedded in the model, and the stress real-time monitoring is realized by a YJZ-3A intelligent digital strain gauge. In each similar material model, the seams are arranged with 7 lines and 8 displacement points. The transverse spacing is 00 mm and the longitudinal spacing is 00 mm. The stress monitoring points are arranged in the roof rock strata 8 m away from the coal seam. The monitoring point spacing is 00 mm, and a total of stress monitoring points are arranged. 4. Similar experiment results and analysis 4.. Overlying strata and slope 3D optical displacement field The XJTUDP 3D optical measurement system was first used herein to obtain photographs and determine the monitoring point displacement of the overlying strata during an underground Figure. Three-dimensional optical photogrammetry system. mining operation (Zhang et al 0). The initial displacement of the surrounding rock was zero. The optical measurement system was used at each stage of the excavation to obtain model photos. The data were then processed using the 3D optical static deformation measurement software. The displacements of all monitoring points on the entire model were calculated based on the differences between all the stages and the original state. oreover, the displacement cloud images were generated. Note that, compared with manual measurement, using this system omits a large amount of tedious work and yields improved measurement accuracy. Figure 3 shows the identification process used for the displacement monitoring points in this experiment. 4.. Slope and overlying strata dynamic movement law The upper work surface was mined first in the examined mining process. A three-belt distribution had then formed at the top of the goaf when the mining was finished. This distribution was characterized by the maximum settlement generated at the top of the mined-out area (Hao et al 0, Wang et al 0). In addition, the collapse of the strata above the goaf was more significant, and a downward bending deformation occurred. Note that the deformation and the movement of the rock were basically identical. The deformation and the movement of the strata above the goaf were more obvious when the lower work face was mined. The settlement of the surface rock above the goaf increased under the influence of the upper work face. oreover, the overlying strata were bent toward the lower region. This behavior produced a horizontal deformation in the surface direction. Figure 4 shows the motion vectors of each measuring point, where area A represents the surface layer, the area above the goaf is labeled B, and C represents the work face sides. The observation point movement vectors pointed to the mined area after the work face mining was finished because of the influence of the underground engineering excavation. Furthermore, the movement gradually increased from the edge of the surface to the goaf top. The roof rock was bent and deformed after the work face was mined (figure 4(b)). The roof degraded, and a fractured rock was formed when the tensile deformation exceeded the ultimate tensile rock strength. The volume of the filling rock gradually increased, 64

6 Figure. Physical map of the model displacement measurement point monitoring area. (a) odel size (m); (b) measuring points; (c) physical model. 643

7 Figure 3. Displaced monitoring point recognition process. (a) odel state before excavation; (b) model state after excavation; (c) calculation results. 644

8 Figure 4. otion vectors of each measuring point in the rock strata. (a) otion vectors of each measuring point over the entire monitoring area; (b) motion vectors of the measuring points above the goaf area; (c) motion vectors of each measuring point in the C region. and the upper-rock movement gradually reduced. As regards the work face sides (figure 4(c)), the observation point moving vectors pointed in the C direction and the movement gradually increased because of the influence of the mining stress and compression deformation (Zhu et al 00, Baharuddin et al 06). The degree of the rock mass movement around the floor was small, while the upper-rock movement was greater. 5. A eshless numerical model for particle contact based on similar experimental laws 5.. Principles of PC Contact pair searching and contact force calculation are two main steps in DE. Due to the abundant information in contact pairs, the elements which are used to calculate deformation force could be formed easily. PC is a method which uses the contact pairs of DE to generate the elements and uses the elements to simulate engineering problems. By renewing neighbors for each particle in each step, the elements will be created or deleted correspondingly, and then the element distortion problems will be solved automatically. Figure 5 shows the main idea of PC. The incrementalbased explicit method is adopted for PC, and forwarddifference approximation is used. In PC, all the field variables (i.e. stress, strain, density, acceleration, velocity and displacement) are stored in the particles. The particles are adopted to compute movement, and the elements are used to calculate deformation force. Several steps should be repeated in each time step: () Neighbor deletion: delete neighbors which do not make contact with the particle. () Neighbor searching: find neighbors for each particle. (3) Element deletion: delete elements which do not satisfy the corresponding condition. 645

9 Figure 5. ain idea of PC. (a) Element creation based on particle contacts; (b) element deletion; (c) element recreation. (4) Element creation: create elements based on new neighbor relationship. (5) Deformation force calculation: based on continuous constitutive law, calculate stress and node force. (6) Contact force calculation: if a particle block contact exists, calculate the contact force. (7) Particle evolvement calculation: based on Newton s law, calculate the movement of each particle. 5.. Element stress and deformation force calculation Incremental-based FV is adopted (Jing and Stephansson 007) to calculate element stress and deformation force. In numerical time t 0, the particle velocity could be obtained from Newton s law, and the velocity gradient of each triangle element could be obtained by Gauss s theorem (equation (8)), where á vi xjñ means the average velocity gradient of the element, S e denotes the volume of the element in D, v i represents the average velocity in edge k, n k j is the jth component of the unit normal vector in edge k and ΔL k is the length of edge k: vi vi x» x j j 3 = å vn i j k DL k. ( 8) S e k= Based on equation (8), the incremental strain of the element could be calculated (equation (9)), where Deij denotes the element incremental strain, and Δt means time step: vi vj D eij = t. 9 x + D ( ) j xi denotes the bulk modulus, G represents the shear modulus, Δθ represents the incremental bulk strain, dij is the Kronecker delta and s ij-old is the element stress in the last step. Δθ could be obtained by equation (), and s ij-old could be calculated p p p3 by equation (3), where sij - old, s ij-old and s ij-old are the stresses of three particles in the last step: D sij = GD eij + K - G Dqdij 3 ( 0) sij =D sij + sij - old ( ) D q = D e + D e + D e33 ( ) s p = ( s p + s p3 + s )/ 3. ( 3) ij-old ij-old ij-old ij-old If some plastic constitutive law is adopted, the stress will be corrected. In this paper, maximum tensile criteria and ohr Coulomb criteria are adopted (equation (4)), where sij - new is the corrected element stress in this step, and T, C and f denote the tensile strength, cohesion and inner friction angle: sij- new = f ( sij, T, C, f). ( 4) According to corrected stress tenser sij - new, the deformation force of the element could be calculated by equation (5), where F p i means the ith component of the deformation force of node P (particle P) in the triangle element (each node owns two corresponding edges): å F = s ( n DL k /. ) ( 5) i p ij-new k= j k According to equations (0) and (), the stress increment Dsij and the stress sij could be obtained, where K Due to the superposition algorithm described in section.3, if a particle cluster contains N particles, the 646

10 Figure 6. Total displacement in the different mining stages. (a) Before the mining; (b) working face distance from a slope toe of 470 m (initial stop line); (c) working face distance from a slope toe of 399 m (reasonable stop line ); (d) working face distance from a slope toe of 34 m (reasonable stop line ); (e) working face distance from a slope toe of 33 m (designed stop line); (f) working face distance from a slope toe of 03 m (continue mining for 60 0 m). contributions of the related elements should be divided by N (equation (6)): F = s - å ( nj k DL k /. ) ( 6) N i p ij new k= After all the stresses and deformation forces of the elements have been calculated, the element strains and stresses should be transformed to particles (equations (7) and (8)), p p where e ij-new and s ij-new denote particle strain and stress in p this step, eij - old means particle strain in the last step, Deij k - new k and sij - new represent the kth element strain increment and stress in this step, and represents the element number related to the particle: 6. Analysis of movement and failure law of slope based on the PCC numerical model 6.. Simulation model In this method, the continuous media elements are created based on the contact topology of particle DE, and the elements will be deleted or recreated according to the movement and evolvement of the particle system (Xu et al 06). The DE s mesh is generated by GSH software. The particles of the PC are generated by the nodes of the DE mesh. The model size is 5000 mm 300 mm 000 mm, with the bottom boundary totally fixed and the side boundary vertically fixed, and the material parameters are as shown in table. p ij-new å e = D e - + e - ( 7) s p ij-new k= k ij new å p ij oid = Deij- new. ( 8) k= k 6.. Slope and overlying strata movement law Figures 6 and 7 show the following mining process: first, the direct roof collapses with a collapse angle of approximately 54 ; the main roof rock will also collapse when it cannot bear the overlying strata force action (the collapse height is approximately m) (Abd 05). The overlying strata then 647

11 increased when the mining face advanced from reasonable stop line to reasonable stop line. Furthermore, the horizontal displacement monitored by measuring point 4 slightly reduced. No significant change in the displacement of measuring point 3 was found during the process of reasonable stop line mining to the designed stop line. The relative displacement of monitoring points 4, 5, 7, and 8 was larger when mined to reasonable stop line. eanwhile, the relative displacement of monitoring points, 5, 7, and 8 was larger when mined to reasonable stop line Failure process of slope and overlying strata Figure 7. Slope displacement during underground mining. (a) Vertical displacement of the slope in different mining stages; (b) relationship between the horizontal displacement and the mining stage. exhibit an obvious abscission layer, and the cracks will develop along the working face, eventually leading to a slow sinking of the overlying rock mass. The horizontal displacement of the free slope surface, when the working face is mined to the design stop line, is approximately 4 m (figure 7(e)). The cracks between the first and second steps are connected, and the step may slide out. Cracks will reach the bottom of the slope if mining is continued for 60 m 0 m. The upper steps squeeze out, and the slope slides (figure 6(f)) Influence of the working face mining on the slope movement Figure 7 shows the monitoring displacement of the slope. The vertical displacement of the slope at different mining stages (figure 7(a)) demonstrates that the vertical displacement of the slope gradually increased with the advance of the mining face. The settlement value of monitoring point 5 was the largest, and the subsidence was completed when mining to the initial stop line. Figure 7(b) shows that the horizontal displacement monitored by measuring point 3 was drastically Figure 8 shows the failure state of the interfacial spring at different mining stages. The overlying rock mass exhibited tensile and shear failure because of the underground mining (Gu and Ozbay 04, Joko et al 07). The slope structure also showed serious damage (Song et al 0, Behrooz et al 07). The upper and lower coal seams were destroyed when the lower coal seam was mined to the initial stop line because of the shearing action (red line in figure 8). The slope failure denoted that the excavation led to a rock mass collapse. In addition, the slope near monitoring points 3 5 presented tension cracks. Therefore, a significant safety hazard, which should be appropriately monitored, is present. 7. Conclusions In this study, a simulation experiment for underground mining is performed. The movement of the overlying strata caused by mining during the mining process is observed using a 3D optical displacement monitoring system. Based on the simulation experiment results, a PC is proposed. The overlying strata and slope deformation process was analyzed using the PC numerical simulation method. The following conclusions can be drawn:. The overlying strata gradually collapsed with the continuous progress of the work face. Furthermore, the basic roof bore the force generated by the overlying rock mass. Hence, a rock beam force distribution formed. In addition, the upper load of the rock beam gradually increased with the increased goaf area, and the stress of the rock beam center was at its largest. Therefore, the rock beam was prone to fracturing at the center. oreover, the collapsed rock layers became superimposed upon each other as the rock strata gradually fractured. At this stage, the monitoring data showed fewer cracks at the rock center above the goaf area and more at the bottom and top of the rock.. The rock strata subsided and induced the beam structure effect when the work face advanced to the design stop line and continued to advance, causing a rock mass partial settlement on the slope and a serious free-face horizontal movement. This behavior influenced the slope stability and could even induce cracks and potential slip and sliding surfaces. The slope was 648

12 Figure 8. Slope and overlying strata failure process. (Note that each point represents a finite element node. The blue node stands for undamaged ; the red node indicates shearing failure ; and the green node indicates tensile failure.) prone to lateral or traction landslides under such conditions. 3. The numerical simulation results indicated that the deformation process of the overlying strata was as follows: the roof immediately starts to collapse the rock mass of the main roof collapses overlying strata generate an abscission layer a fracture of the free surface develops overlying strata begins to sink. eanwhile, the deformation and failure process of the slope is as follows: overlying rock mass exhibits shear failure the overburden is completely sheared the overburden collapses the slope surface exhibits a tensile crack the stepped crack is connected the slope begins to slide to the outside. 4. Similar simulation experiment results of the slope deformation and failure laws are consistent with the numerical simulation results, which proved that the PC numerical calculation method is feasible, providing a new simulation method for slope elastoplastic analysis and sliding process monitoring. Acknowledgments This work is supported by the Special Funds for Central Guiding Local Science and Technology Development (project number: Z ), National Natural Science Foundation of China (project number: , ), Beijing Nova Program (project number: Z ). The author is grateful for the comments and suggestions provided by the editors and reviewers, which helped to improve the quality of this paper. References Abd A H 05 Earthquake-induced displacement of cohesivefrictional slopes subject to cracks IOP Conf. Ser.: Earth Environ. Sci Baharuddin I N Z, Omar R C, Roslan R, Khalid N H N and Hanifah I 06 Determination of slope instability using spatially integrated mapping framework IOP Conf. Ser.: ater. Sci. Eng Behrooz G, Ren G, John S and Lucas H 05 Application of 3D laser scanner, optical transducers and digital image processing 649

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