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Stress redistribution of longwall mining stope and gas control of multi-layer coal seams
International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
Contents lists available at ScienceDirect
International Journal of Rock Mechanics & Mining Sciences journal homepage: www.elsevier.com/locate/ijrmms
Technical Note
Stress redistribution of longwall mining stope and gas control of multi-layer coal seams Wei Yang a,b,n, Bai-quan Lin a,c, Qing Yan d, Cheng Zhai a,c a
Faculty of Safety Engineering, China University of Mining & Technology, Xuzhou, Jiangsu, China Lu'an Mining Group, Changzhi, Shanxi, China c Key Laboratory of Gas and Fire Control for Coal Mines, China University of Mining & Technology, Xuzhou, Jiangsu, China d Lu'an Group Gaohe Energy Co. Ltd., Shanxi 047100, China b
art ic l e i nf o Article history: Received 27 November 2013 Received in revised form 23 May 2014 Accepted 21 August 2014
1. Introduction The longwall mining method is widely used all over the world. Underground coal mining has improved both from production and productivity with the development of the mechanized longwall mining method, but there are still certain risks that can result in unacceptable levels of safety. When there are multi-layers of coal seams, one of the most dangerous issues is the pressure-relieved gas from the adjacent coal seams, which may cause a gas explosion. During the mining process, the three-dimensional stress distribution of the stope will change and cause the fractures to open or close, and then the adsorption gas in the adjacent coal seam begins to flow to the coalface and gob along the fractures by the pressure gradient of itself [1–3]. Stress redistribution around longwall coalface was widely studied mainly focus on the roof fall and ground control [4,5]. The passive seismic velocity tomography and geostatistical estimation were used to investigate the stress redistribution around the longwall mining panel [6,7]. The ground behavior, reinforcement performance and stress redistribution was studied when the longwall panel subjected to a severe horizontal stress concentration [8]. The vertical stress changes in multi-seam mining under supercritical longwall panels has also been studied [9]. Numerical simulation method was used to study the longwall mining-induced strata movement, stress and fracture evolution of the mining stope, and gas flow characteristics in underground coal mine, which is useful for the coal mine ventilation system design and gas drainage [10–16]. A three-dimensional annular-shaped
n Corresponding author at: Faculty of Safety Engineering, China University of Mining & Technology, Xuzhou, Jiangsu 221116, China. E-mail address: [email protected] (W. Yang).
http://dx.doi.org/10.1016/j.ijrmms.2014.08.009 1365-1609/& 2014 Elsevier Ltd. All rights reserved.
stress relief zone along the perimeter of the longwall panel was identified for optimal methane drainage [17]. Some relative permeability and stress-damage-flow coupling models for coal were built, and it was found that the relative permeability change can significantly affect the gas production rate [18–20], and the permeability change with the swelling was also studied [21]. The previous studies are of great help for coal mine support and gas control, but the engineers may still has confusion for the pressure relief gas control, for there are still certain problems unsolved. For example, how the stress redistribution affected by the mining height, how the stress changes during the mining process and how the stress evolution affect gas flow have been little researched. This paper systematically investigated the mining stope evolution and division in a longwall mining panel to solve these problems for the pressure relief gas control.
2. Investigation method In order to study the stress distribution and evolution under different mining height in a longwall panel, and then to study the pressure relief gas flow characteristics, this paper used numerical simulation software FLAC3D for auxiliary analysis. FLAC3D is a three-dimensional explicit finite-difference program for engineering mechanics computation. The explicit, lagrangian, calculation scheme and the mixed-discretization zoning technique used in FLAC3D ensure that plastic collapse and flow are modeled very accurately. Before the final model was built, several simple models were built to test the boundary effect of model size and boundary conditions. The final model mesh grids, size and boundary are shown in Fig. 1.
W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
9
Fig. 1. Numerical model of 3D view (a) model mesh grids and (b) model size and boundary.
The final three-layered model including roof, coal seam and floor has 405108 zones and 426321 grid-points with the 3D of 900 m along the y-direction, 750 m along the x-direction and 744 m along the z-direction. The coal seam thickness is 6 m with dip angle of 01, roof thickness of 394 m, floor thickness of 344 m. The panel is 150 m width in the x-direction, 300 m length in the y-direction, and the panel is in the middle of the model with 300 m of surrounding rock. The mining height is between 1 m and 6 m. At the same grid of the model, the initial stresses are equal in all directions, and the average density is 2500 kg/m3. Compressive initial stress of 10 MPa is applied on the top of the model, the initial stress is 20 MPa at the mined coal seam level and 28.6 MPa at the bottom calculated by the acceleration of gravity and rock mass density. During the simulating process, the elastic-plastic model and the Mohr–Coulomb failure criterion were applied to the threelayered model, which can represent shear failure in rocks and is widely used in underground engineering. Rock mass and coal seam undergo plastic flow when their yield limits are reached. The roof and the floor mechanical properties were assumed the same. The mechanical properties of different rock layers are shown in Table 1.
The panel mined from the start position to the end position along the y-direction, the excavation length is 5 m every cycle, the next cycle will not excavate until the model get to a new balanced state. The gob was not filled, and the roof and floor in the gob should de-stress and undergo plastic flow and get in touch with each other about 10 m behind the coalface. The model simulates six times at different mining heights of 1–6 m to investigate how the stress redistribution affected by the mining height.
3. Results 3.1. Stress redistribution along z-direction at different mining height First, we analyzed the three-dimensional stress redistribution along the vertical direction when the mining height is 2 m. Fig. 2 shows the stress redistribution of the location 200 m behind the coalface in the middle of the gob. The positive value of the ordinate means above the mined coal seam, negative value means below the coal seam. The negative value of the abscissa means compressive stress. The lines of original and 80% of the original stress are in fig. 2. In this paper, the area on the left side of the line
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W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
of 80% of the original stress is called the stress-decreased part, the area on the right side of the line of original stress is called the stress-increased part, and the area between the lines of original and 80% of original stress is called the transition part. Fig. 2 shows that the three-dimensional stresses along vertical direction of the panel have significant difference, and it mainly represent that the stress drop much more significantly near the coal seam than far away. At the same location, the z-stress is the minimum and the x-stress is the maximum. All of the threedimensional stresses are less than the 80% of the original stress from 54 m in floor to 79 m in roof; three-dimensional stresses in this range are reduced significantly, so this range was called the total de-stressed belt. Only z-stress is less than the 80% of the original stress in the range of 79–185 m in roof and 54 m to 149 m in floor, the horizontal stresses are still high, and the x-stress even increased, these two ranges are called vertical destressed belt. The three-dimensional stresses had no significant change beyond the 185 m in roof and 149 m in floor, and the range is called the original belt. The division of the belts according to the characteristics of the stress distribution is under the condition of 2 m mining height. In order to study how the stress relief range affected by the mining height, the stress distributions were studied when the mining height was 1–6 m separately. The vertical stress distributions of the different mining height are shown in Fig. 3.
Table 1 Model parameters Parameters Density (kg/m3)
Bulk modulus (GPa)
Shear modulus (GPa)
Friction Cohesion angle (1) (MPa)
Tension (MPa)
Coal seam Roof Floor
1 2 2
0.8 1.6 1.6
20 25 25
0.5 1 1
1450 2500 2500
0.5 1 1
According to the belt division method like Fig. 2, the upper and lower bounds of total de-stressed belt and the upper and lower bounds of vertical de-stressed belt at different mining height are shown in Fig. 4. The mining height significantly affects the de-stressed range, the de-stressed range increases with the increasing mining height, but not following the constant proportional. Generally, the destressed range in roof is larger than in floor.
3.2. Stress evolution during the mining process We already analyzed the three-dimensional stress distribution along the z-direction at the location 200 m behind the coalface in the middle of the gob in Section 3.1, and the stress evolution process during the mining process of the specific location will be analyzed in this part. The locations of the test points are shown in Fig. 5, and the mining height is 2 m. During the entire mining process, the maximum and minimum principle stresses of the test points were consistently recorded, and the stresses changing with the coalface location are plotted in Fig. 6(a) and (b) separately. Fig. 6(a) shows that no minimum principal stresses of the test points increase in the entire mining process. The nearer apart from the coal seam, the more significantly decrease the minimum principal stress. The minimum principal stresses of the test points between 150 m in floor and 150 m in roof reduce significantly, while the minimum principal stresses of the test points beyond 200 m in floor and 200 m in roof remains nearly unchanged throughout the production process. Fig. 6(a) shows that the maximum principal stresses of the test points between 50 m in floor and 50 m in roof increase before the coalface and decrease behind the coalface, which suggests these test points in the total de-stressed belt. The maximum principal stresses of the test points of 100 m, 150 m in floor and 100 m, 150 m in roof increase behind the coalface, which suggests these test points in the vertical de-stressed belt. Beyond 200 m in floor and 200 m in roof, the maximum principal stress
Fig. 2. Three-dimensional stress distribution along z-direction in the gob. Note: x-stress is the stress along the x-direction, y-stress is the stress along the y-direction, and the z-stress is the stress along the z-direction.
W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
1m
2m
3m
4m
5m
6m
11
Fig. 3. Principal stress distribution along z-direction at different mining height.
remains nearly unchanged, which suggests these test points in the original belt. By comparing Fig. 6(a) and (b), it can be seen that the stope stress change dynamically in the mining process. The different principal stresses of the same test points may have very different
changing trends. The rock mass 50 m to the coal seam particularly experiences the complex stress evolution, which will greatly increase the fractures mutual connection and then increase the permeability. The different change trends of the maximum and minimum principle stresses of the rock mass between 100 m and
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W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
Fig. 4. The range of the different belts at different mining heights.
-300 -250 -200 -150
end mining position
floor
300 m
20 50 100
y-direction mining direction
100 m
4. Discussion and application start mining position
-100
-50 -10
coal seam
-20
10
300 250 200 150
roof
decreased. In the stress-increased circle, the maximum principle stress increase, and Fig. 7(b) and (d) shows that the maximum principle stress is along the tangential direction of the stress-increased circle, the stress perpendicular to the tangential direction of the stressincreased circle decreases. Stress in the original zone remains nearly unchanged. Massive experiments focus on relationship between permeability and stress of fractural rock mass generally proved that the permeability would increase when the fractures expand with the reduced normal stress, and the permeability will decrease when the fractures close with the increased normal stress [24–30]. The shear or dilation of the fracture may cause complex deformation of the rock, but the fracture deformation subjects to the normal stress is usually much more important, so in the paper the shear or dilation of fractures was neglected. In the stress decreased circle, stresses along every direction reduce, the rock swell, the internal fractures propagate and mutual communicate to make good gas flow channels, which will greatly improve the permeability. Gas in this region can flow freely from the high-pressure region to the low-pressure region by the gradient of itself. In the stress-increased circle, fractures perpendicular to the tangential of the stressincreased circle will close with the increased tangential stress, which will cause the gas flowing through the stress-increased circle more difficult. However, the fractures along the tangential of the stress-increased circle will expand, and the gas is much more easily to flow along the circle. The different gas flow characteristics in different circles may sufficiently serve for gas control.
0m
Fig. 5. Location of the test points.
150 m to the coal seam will cause the permeability different in different directions [22,23], which will significantly affect the gas flowing characteristics. 3.3. Division of the mining stope The complex stress variation of the stope has a significant influence on gas control and roof support. Fig. 7(a) and (c) shows the distribution of maximum principle stress value on the strike and dip profiles. Fig. 7(b) and (d) shows the principal stress direction of each unit on the strike and dip profiles, the line length in each unit indicates the principle stress value, and the direction of the line represent the principle stress direction. Generally, the principal stress value and direction change greatly around the gob. Fig. 7(a) and (c) shows that the maximum principal stress redistribution around the mined out area can be divided into three concentric circles both on the strike and dip profile including stress-decreased circle, stress-increased circle and the original zone. The three concentric circles correspond to the three belts of total de-stressed belt, vertical de-stressed belt and original belt divided in Section 3.1. In the stress-decreased circle, the maximum principle stress is lower than the original stress, that is to say stresses in every direction
In the condition of multi-layer of coal seams, when the first coal seam is mined, the three-dimensional stresses of adjacent coal seams in the stress-decreased circle would decrease. Gas from the upper coal seams would flow downward into the porous stress decreased circle in roof by the gas pressure gradient. In addition, the caving-in may compress the storage gas in roof into the coalface and cause the gas overrun. Gas from the lower coal seams would flow upward through the stress-decreased circle in floor and into the gob and stressdecreased circle in roof. At the same time, the coalface ventilation will affect the gas upward flowing route in some extent. Some of the air would leak into the gob when the air ventilates from the inlet roadway to the outlet roadway. The leaked air will bring some high concentration gas in the gob into the upper corner near the air outlet roadway, which may cause the gas concentration into the explosive range. Above all, we can find three gas enrichment zones including the gas upward flow channel of the stress-decreased circle in floor, the stress-decreased circle in roof and the upper corner near the air outlet roadway. According to the gas flow and enrichment behaviors, the comprehensive gas control technology is shown in Fig. 8. Downward boreholes from air outlet roadway or some other adjacent roadways to the stress-decreased zone in floor should be set to drain the upward flowing gas to change the gas flowing direction and to prevent the gas from flowing upward. Upward boreholes from air outlet roadway or some other adjacent roadways to the stressdecreased circle in roof should be set to drain the enrichment gas. Consider the traditional U-shaped ventilation method would bring the gas in the gob into the upper corner and cause gas overrun, a new Y-shaped ventilation method was developed. Based on the U-shaped ventilation method, the air outlet roadway is preserved by filling and supporting technologies after the coalface in the gob as the tail roadway for special return airway, and the traditional air outlet roadway of the U-shaped ventilation method is changed into the air inlet roadway, which is shown in Fig. 9. Thus, there are two roadways for air inlet and one preserved tail roadway for air outlet. All of the
W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
13
Fig. 6. Principle stresses evolution of the test points during the mining process (a) minimum principle stress and (b) maximum principle stress.
machines and workers should be in the air inlet roadways to keep security. In the filling area, air vents should be set every 5–7 m to connect the gob and tail roadway to form the ventilation system in the gob, to change the airflow route in the gob to prevent the upper corner gas overrun. When the adjacent coal seam is in the stress increased circle, the gas desorption can still occur, nevertheless the gas can much more freely flow along the circle rather than through. The desorption gas can be drained by the boreholes, otherwise the gas adsorption process may restart after the coalface to keep the high gas content in the coal seam, and the gas outburst risk cannot be canceled. When the techniques of gas control used in a specific mine, the geological strata section of the mine must be concerned, for the geological strata can modify the sizes of the zones divided above.
5. Conclusions (1) Stress distribution along the vertical direction was investigated by the numerical simulation method under different mining heights and was divided into three belts including total
de-stressed belt, vertical de-stressed belt and original belt. The stress-decreased ranges increase with the growing mining height, but not following a constant proportion. (2) The stress evolution of different locations was investigated during the entire mining process. The maximum and minimum principal stresses at the same test point may have quite different changing trends. The maximum principle stress changes much more complex than the minimum principle stress, which may cause the permeability at the same location different along different directions. (3) The mining stope was divided into three concentric annuluses including the stress-decreased circle, stress increased circle and the original zone. The gas flow and enrichment characters were analyzed in different circles according to the stress value and direction, based on which, a gas control technology was developed by the combination of the boreholes and ventilation system. (4) It should be noted that there is no interface or joint element between the different layers (roof, coal seam and floor); the rock mass is modeled as an anisotropic media with three layers that have been completely welded together.
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Fig. 7. Maximum principal stress distribution and division of the mining stope (a) stress value in strike profile, (b) stress direction in strike profile, (c) stress value in dip profile, and (d) stress direction in dip profile.
Fig. 8. Comprehensive gas control technology according to gas flow and enrichment character.
W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
15
Fig. 9. Y-shaped ventilation system.
Acknowledgments This work was supported by the Fundamental Research Funds for the Central Universities (2014QNA02), Program for Changjiang Scholars and Innovative Research Team in University (Grant no. IRT13098), China Postdoctoral Science Foundation funded Project (2014M551057), National Natural Science Foundation of China (51404261), and Natural Science Foundation of Jiangsu province (BK20140196). References [1] Yang W, Lin B, Qu Y, Zhao S, Zhai C, Jia L, et al. Mechanism of strata deformation under protective seam and its application for relieved methane control. Int J Coal Geol 2011;85:300–6. [2] Li J, Su X, Na M, Yang H, Li J, Yu Y, et al. Influence of gas flow on thermal field and stress during growth of sapphire single crystal using the Kyropoulos method. Rare Metals 2006;25:260–6. [3] Wang JG, Kabir A, Liu JS, Chen ZW. Effects of non-Darcy flow on the performance of coal seam gas wells. Int J Coal Geol 2012;93:62–74. [4] Richards MJ. Longwall front abutment stress effects firedamp release. Min Sci Tech 1984;1:215–29. [5] Yang W, Lin B, Qu Y, Li Z, Zhai C, Jia L, et al. Stress evolution with time and space during mining of a coal seam. Int J Rock Mech Min Sci 2011;48:1145–52. [6] Hosseini N, Oraee K, Shahriar K, Goshtasbi K. Studying the stress redistribution around the longwall mining panel using passive seismic velocity tomography and geostatistical estimation. Arab J Geosci 2013;6:1407–16. [7] Luxbacher K, Westman E, Swanson P, Karfakis M. Three-dimensional timelapse velocity tomography of an underground longwall panel. Int J Rock Mech Min Sci 2008;45:478–85. [8] Mark C, Gale W, Oyler D, Chen J. Case history of the response of a longwall entry subjected to concentrated horizontal stress. Int J Rock Mech Min Sci 2007;44:210–21. [9] Suchowerska AM, Merifield RS, Carter JP. Vertical stress changes in multi-seam mining under supercritical longwall panels. Int J Rock Mech Min Sci 2013;61:306–20. [10] Islam MR, Shinjo R. Numerical simulation of stress distributions and displacements around an entry roadway with igneous intrusion and potential sources of seam gas emission of the Barapukuria coal mine, NW Bangladesh. Int J Coal Geol 2009;78:249–62. [11] Zhou H, Liu J, Xue D, Yi H, Xue J. Numerical simulation of gas flow process in mining-induced crack network. Int J Min Sci Tech 2012;22:793–9. [12] Whittles DN, Lowndes IS, Kingman SW, Yates C, Jobling S. Influence of geotechnical factors on gas flow experienced in a UK longwall coal mine panel. Int J Rock Mech Min Sci 2006;43:369–87.
[13] Islam MR, Hayashi D. Kamruzzaman ABM. Finite element modeling of stress distributions and problems for multi-slice longwall mining in Bangladesh, with special reference to the Barapukuria coal mine. Int J Rock Mech Min Sci 2009;78:91–109. [14] Nawrocki PA, Mróz Z. A constitutive model for rock accounting for viscosity and yield stress degradation. Comput Geotech 1999;25:247–80. [15] Shabanimashcool M, Li CC. A numerical study of stress changes in barrier pillars and a border area in a longwall coal mine. Int J Coal Geol 2013;106:39–47. [16] Wang HW, Jiang YD, Zhao YX, Zhu J, Liu S. Numerical investigation of the dynamic mechanical state of a coal pillar during longwall mining panel extraction. Rock Mech Rock Eng 2013;46:1211–21. [17] Guo H, Yuan L, Shen B, Qu Q, Xue J. Mining-induced strata stress changes, fractures and gas flow dynamics in multi-seam longwall mining. Int J Rock Mech Min Sci 2012;54:129–39. [18] Chen D, Pan Z, Liu J, Connell LD. An improved relative permeability model for coal reservoirs. Int J Coal Geol 2013;109–110:45–57. [19] Gentzis T. Economic coalbed methane production in the Canadian Foothills: solving the puzzle. Int J Coal Geol 2006;65:79–92. [20] Yang TH, Xu T, Liu HY, Tang CA, Shi BM, Yu QX. Stress–damage–flow coupling model and its application to pressure relief coal bed methane in deep coal seam. Int J Coal Geol 2011;86:357–66. [21] Wu Y, Liu J, Elsworth D, Siriwardane H, Miao X. Evolution of coal permeability: contribution of heterogeneous swelling processes. Int J Coal Geol 2011;88:152–62. [22] Wang JG, Liu JS, Kabir A. Combined effects of directional compaction, non-Darcy flow and anisotropic swelling on coal seam gas extraction. Int J Coal Geol 2013;109:1–14. [23] Yang W, Lin BQ, Zhai C, Li XZ, An S. How in situ stresses and the driving cycle footage affect the gas outburst risk of driving coal mine roadway. Tunn Undergr Sp Tech 2012;31:139–48. [24] Bandis SC, Lumsden AC, Barton NR. Fundamentals of rock joint deformation. Int J Rock Mech Min Sci 1983;20:249–68. [25] Bandis SC, Lumsden AC, Barton NR. Experimental studies of scale effects on the shear behaviour of rock joints. Int J Rock Mech Min Sci 1981;18:1–21. [26] Baghbanan A, Jing L. Stress effects on permeability in a fractured rock mass with correlated fracture length and aperture. Int J Rock Mech Min 2008;45:1320–34. [27] Chen M, Bai M. Modeling stress-dependent permeability for anisotropic fractured porous rocks. Int J Rock Mech Min Sci 1998;35:1113–9. [28] He J, Chen S, Shahrour I. Numerical estimation and prediction of stressdependent permeability tensor for fractured rock masses. Int J Rock Mech Min Sci 2013;59:70–9. [29] Latham J, Xiang J, Belayneh M, Nick HM, Tsang C, Blunt MJ. Modeling stressdependent permeability in fractured rock including effects of propagating and bending fractures. Int J Rock Mech Min Sci 2013;57:100–12. [30] Min K, Rutqvist J, Tsang C, Jing L. Stress-dependent permeability of fractured rock masses: a numerical study. Int J Rock Mech Min Sci 2004;41:1191–210.
Contents lists available at ScienceDirect
International Journal of Rock Mechanics & Mining Sciences journal homepage: www.elsevier.com/locate/ijrmms
Technical Note
Stress redistribution of longwall mining stope and gas control of multi-layer coal seams Wei Yang a,b,n, Bai-quan Lin a,c, Qing Yan d, Cheng Zhai a,c a
Faculty of Safety Engineering, China University of Mining & Technology, Xuzhou, Jiangsu, China Lu'an Mining Group, Changzhi, Shanxi, China c Key Laboratory of Gas and Fire Control for Coal Mines, China University of Mining & Technology, Xuzhou, Jiangsu, China d Lu'an Group Gaohe Energy Co. Ltd., Shanxi 047100, China b
art ic l e i nf o Article history: Received 27 November 2013 Received in revised form 23 May 2014 Accepted 21 August 2014
1. Introduction The longwall mining method is widely used all over the world. Underground coal mining has improved both from production and productivity with the development of the mechanized longwall mining method, but there are still certain risks that can result in unacceptable levels of safety. When there are multi-layers of coal seams, one of the most dangerous issues is the pressure-relieved gas from the adjacent coal seams, which may cause a gas explosion. During the mining process, the three-dimensional stress distribution of the stope will change and cause the fractures to open or close, and then the adsorption gas in the adjacent coal seam begins to flow to the coalface and gob along the fractures by the pressure gradient of itself [1–3]. Stress redistribution around longwall coalface was widely studied mainly focus on the roof fall and ground control [4,5]. The passive seismic velocity tomography and geostatistical estimation were used to investigate the stress redistribution around the longwall mining panel [6,7]. The ground behavior, reinforcement performance and stress redistribution was studied when the longwall panel subjected to a severe horizontal stress concentration [8]. The vertical stress changes in multi-seam mining under supercritical longwall panels has also been studied [9]. Numerical simulation method was used to study the longwall mining-induced strata movement, stress and fracture evolution of the mining stope, and gas flow characteristics in underground coal mine, which is useful for the coal mine ventilation system design and gas drainage [10–16]. A three-dimensional annular-shaped
n Corresponding author at: Faculty of Safety Engineering, China University of Mining & Technology, Xuzhou, Jiangsu 221116, China. E-mail address: [email protected] (W. Yang).
http://dx.doi.org/10.1016/j.ijrmms.2014.08.009 1365-1609/& 2014 Elsevier Ltd. All rights reserved.
stress relief zone along the perimeter of the longwall panel was identified for optimal methane drainage [17]. Some relative permeability and stress-damage-flow coupling models for coal were built, and it was found that the relative permeability change can significantly affect the gas production rate [18–20], and the permeability change with the swelling was also studied [21]. The previous studies are of great help for coal mine support and gas control, but the engineers may still has confusion for the pressure relief gas control, for there are still certain problems unsolved. For example, how the stress redistribution affected by the mining height, how the stress changes during the mining process and how the stress evolution affect gas flow have been little researched. This paper systematically investigated the mining stope evolution and division in a longwall mining panel to solve these problems for the pressure relief gas control.
2. Investigation method In order to study the stress distribution and evolution under different mining height in a longwall panel, and then to study the pressure relief gas flow characteristics, this paper used numerical simulation software FLAC3D for auxiliary analysis. FLAC3D is a three-dimensional explicit finite-difference program for engineering mechanics computation. The explicit, lagrangian, calculation scheme and the mixed-discretization zoning technique used in FLAC3D ensure that plastic collapse and flow are modeled very accurately. Before the final model was built, several simple models were built to test the boundary effect of model size and boundary conditions. The final model mesh grids, size and boundary are shown in Fig. 1.
W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
9
Fig. 1. Numerical model of 3D view (a) model mesh grids and (b) model size and boundary.
The final three-layered model including roof, coal seam and floor has 405108 zones and 426321 grid-points with the 3D of 900 m along the y-direction, 750 m along the x-direction and 744 m along the z-direction. The coal seam thickness is 6 m with dip angle of 01, roof thickness of 394 m, floor thickness of 344 m. The panel is 150 m width in the x-direction, 300 m length in the y-direction, and the panel is in the middle of the model with 300 m of surrounding rock. The mining height is between 1 m and 6 m. At the same grid of the model, the initial stresses are equal in all directions, and the average density is 2500 kg/m3. Compressive initial stress of 10 MPa is applied on the top of the model, the initial stress is 20 MPa at the mined coal seam level and 28.6 MPa at the bottom calculated by the acceleration of gravity and rock mass density. During the simulating process, the elastic-plastic model and the Mohr–Coulomb failure criterion were applied to the threelayered model, which can represent shear failure in rocks and is widely used in underground engineering. Rock mass and coal seam undergo plastic flow when their yield limits are reached. The roof and the floor mechanical properties were assumed the same. The mechanical properties of different rock layers are shown in Table 1.
The panel mined from the start position to the end position along the y-direction, the excavation length is 5 m every cycle, the next cycle will not excavate until the model get to a new balanced state. The gob was not filled, and the roof and floor in the gob should de-stress and undergo plastic flow and get in touch with each other about 10 m behind the coalface. The model simulates six times at different mining heights of 1–6 m to investigate how the stress redistribution affected by the mining height.
3. Results 3.1. Stress redistribution along z-direction at different mining height First, we analyzed the three-dimensional stress redistribution along the vertical direction when the mining height is 2 m. Fig. 2 shows the stress redistribution of the location 200 m behind the coalface in the middle of the gob. The positive value of the ordinate means above the mined coal seam, negative value means below the coal seam. The negative value of the abscissa means compressive stress. The lines of original and 80% of the original stress are in fig. 2. In this paper, the area on the left side of the line
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W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
of 80% of the original stress is called the stress-decreased part, the area on the right side of the line of original stress is called the stress-increased part, and the area between the lines of original and 80% of original stress is called the transition part. Fig. 2 shows that the three-dimensional stresses along vertical direction of the panel have significant difference, and it mainly represent that the stress drop much more significantly near the coal seam than far away. At the same location, the z-stress is the minimum and the x-stress is the maximum. All of the threedimensional stresses are less than the 80% of the original stress from 54 m in floor to 79 m in roof; three-dimensional stresses in this range are reduced significantly, so this range was called the total de-stressed belt. Only z-stress is less than the 80% of the original stress in the range of 79–185 m in roof and 54 m to 149 m in floor, the horizontal stresses are still high, and the x-stress even increased, these two ranges are called vertical destressed belt. The three-dimensional stresses had no significant change beyond the 185 m in roof and 149 m in floor, and the range is called the original belt. The division of the belts according to the characteristics of the stress distribution is under the condition of 2 m mining height. In order to study how the stress relief range affected by the mining height, the stress distributions were studied when the mining height was 1–6 m separately. The vertical stress distributions of the different mining height are shown in Fig. 3.
Table 1 Model parameters Parameters Density (kg/m3)
Bulk modulus (GPa)
Shear modulus (GPa)
Friction Cohesion angle (1) (MPa)
Tension (MPa)
Coal seam Roof Floor
1 2 2
0.8 1.6 1.6
20 25 25
0.5 1 1
1450 2500 2500
0.5 1 1
According to the belt division method like Fig. 2, the upper and lower bounds of total de-stressed belt and the upper and lower bounds of vertical de-stressed belt at different mining height are shown in Fig. 4. The mining height significantly affects the de-stressed range, the de-stressed range increases with the increasing mining height, but not following the constant proportional. Generally, the destressed range in roof is larger than in floor.
3.2. Stress evolution during the mining process We already analyzed the three-dimensional stress distribution along the z-direction at the location 200 m behind the coalface in the middle of the gob in Section 3.1, and the stress evolution process during the mining process of the specific location will be analyzed in this part. The locations of the test points are shown in Fig. 5, and the mining height is 2 m. During the entire mining process, the maximum and minimum principle stresses of the test points were consistently recorded, and the stresses changing with the coalface location are plotted in Fig. 6(a) and (b) separately. Fig. 6(a) shows that no minimum principal stresses of the test points increase in the entire mining process. The nearer apart from the coal seam, the more significantly decrease the minimum principal stress. The minimum principal stresses of the test points between 150 m in floor and 150 m in roof reduce significantly, while the minimum principal stresses of the test points beyond 200 m in floor and 200 m in roof remains nearly unchanged throughout the production process. Fig. 6(a) shows that the maximum principal stresses of the test points between 50 m in floor and 50 m in roof increase before the coalface and decrease behind the coalface, which suggests these test points in the total de-stressed belt. The maximum principal stresses of the test points of 100 m, 150 m in floor and 100 m, 150 m in roof increase behind the coalface, which suggests these test points in the vertical de-stressed belt. Beyond 200 m in floor and 200 m in roof, the maximum principal stress
Fig. 2. Three-dimensional stress distribution along z-direction in the gob. Note: x-stress is the stress along the x-direction, y-stress is the stress along the y-direction, and the z-stress is the stress along the z-direction.
W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
1m
2m
3m
4m
5m
6m
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Fig. 3. Principal stress distribution along z-direction at different mining height.
remains nearly unchanged, which suggests these test points in the original belt. By comparing Fig. 6(a) and (b), it can be seen that the stope stress change dynamically in the mining process. The different principal stresses of the same test points may have very different
changing trends. The rock mass 50 m to the coal seam particularly experiences the complex stress evolution, which will greatly increase the fractures mutual connection and then increase the permeability. The different change trends of the maximum and minimum principle stresses of the rock mass between 100 m and
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W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
Fig. 4. The range of the different belts at different mining heights.
-300 -250 -200 -150
end mining position
floor
300 m
20 50 100
y-direction mining direction
100 m
4. Discussion and application start mining position
-100
-50 -10
coal seam
-20
10
300 250 200 150
roof
decreased. In the stress-increased circle, the maximum principle stress increase, and Fig. 7(b) and (d) shows that the maximum principle stress is along the tangential direction of the stress-increased circle, the stress perpendicular to the tangential direction of the stressincreased circle decreases. Stress in the original zone remains nearly unchanged. Massive experiments focus on relationship between permeability and stress of fractural rock mass generally proved that the permeability would increase when the fractures expand with the reduced normal stress, and the permeability will decrease when the fractures close with the increased normal stress [24–30]. The shear or dilation of the fracture may cause complex deformation of the rock, but the fracture deformation subjects to the normal stress is usually much more important, so in the paper the shear or dilation of fractures was neglected. In the stress decreased circle, stresses along every direction reduce, the rock swell, the internal fractures propagate and mutual communicate to make good gas flow channels, which will greatly improve the permeability. Gas in this region can flow freely from the high-pressure region to the low-pressure region by the gradient of itself. In the stress-increased circle, fractures perpendicular to the tangential of the stressincreased circle will close with the increased tangential stress, which will cause the gas flowing through the stress-increased circle more difficult. However, the fractures along the tangential of the stress-increased circle will expand, and the gas is much more easily to flow along the circle. The different gas flow characteristics in different circles may sufficiently serve for gas control.
0m
Fig. 5. Location of the test points.
150 m to the coal seam will cause the permeability different in different directions [22,23], which will significantly affect the gas flowing characteristics. 3.3. Division of the mining stope The complex stress variation of the stope has a significant influence on gas control and roof support. Fig. 7(a) and (c) shows the distribution of maximum principle stress value on the strike and dip profiles. Fig. 7(b) and (d) shows the principal stress direction of each unit on the strike and dip profiles, the line length in each unit indicates the principle stress value, and the direction of the line represent the principle stress direction. Generally, the principal stress value and direction change greatly around the gob. Fig. 7(a) and (c) shows that the maximum principal stress redistribution around the mined out area can be divided into three concentric circles both on the strike and dip profile including stress-decreased circle, stress-increased circle and the original zone. The three concentric circles correspond to the three belts of total de-stressed belt, vertical de-stressed belt and original belt divided in Section 3.1. In the stress-decreased circle, the maximum principle stress is lower than the original stress, that is to say stresses in every direction
In the condition of multi-layer of coal seams, when the first coal seam is mined, the three-dimensional stresses of adjacent coal seams in the stress-decreased circle would decrease. Gas from the upper coal seams would flow downward into the porous stress decreased circle in roof by the gas pressure gradient. In addition, the caving-in may compress the storage gas in roof into the coalface and cause the gas overrun. Gas from the lower coal seams would flow upward through the stress-decreased circle in floor and into the gob and stressdecreased circle in roof. At the same time, the coalface ventilation will affect the gas upward flowing route in some extent. Some of the air would leak into the gob when the air ventilates from the inlet roadway to the outlet roadway. The leaked air will bring some high concentration gas in the gob into the upper corner near the air outlet roadway, which may cause the gas concentration into the explosive range. Above all, we can find three gas enrichment zones including the gas upward flow channel of the stress-decreased circle in floor, the stress-decreased circle in roof and the upper corner near the air outlet roadway. According to the gas flow and enrichment behaviors, the comprehensive gas control technology is shown in Fig. 8. Downward boreholes from air outlet roadway or some other adjacent roadways to the stress-decreased zone in floor should be set to drain the upward flowing gas to change the gas flowing direction and to prevent the gas from flowing upward. Upward boreholes from air outlet roadway or some other adjacent roadways to the stressdecreased circle in roof should be set to drain the enrichment gas. Consider the traditional U-shaped ventilation method would bring the gas in the gob into the upper corner and cause gas overrun, a new Y-shaped ventilation method was developed. Based on the U-shaped ventilation method, the air outlet roadway is preserved by filling and supporting technologies after the coalface in the gob as the tail roadway for special return airway, and the traditional air outlet roadway of the U-shaped ventilation method is changed into the air inlet roadway, which is shown in Fig. 9. Thus, there are two roadways for air inlet and one preserved tail roadway for air outlet. All of the
W. Yang et al. / International Journal of Rock Mechanics & Mining Sciences 72 (2014) 8–15
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Fig. 6. Principle stresses evolution of the test points during the mining process (a) minimum principle stress and (b) maximum principle stress.
machines and workers should be in the air inlet roadways to keep security. In the filling area, air vents should be set every 5–7 m to connect the gob and tail roadway to form the ventilation system in the gob, to change the airflow route in the gob to prevent the upper corner gas overrun. When the adjacent coal seam is in the stress increased circle, the gas desorption can still occur, nevertheless the gas can much more freely flow along the circle rather than through. The desorption gas can be drained by the boreholes, otherwise the gas adsorption process may restart after the coalface to keep the high gas content in the coal seam, and the gas outburst risk cannot be canceled. When the techniques of gas control used in a specific mine, the geological strata section of the mine must be concerned, for the geological strata can modify the sizes of the zones divided above.
5. Conclusions (1) Stress distribution along the vertical direction was investigated by the numerical simulation method under different mining heights and was divided into three belts including total
de-stressed belt, vertical de-stressed belt and original belt. The stress-decreased ranges increase with the growing mining height, but not following a constant proportion. (2) The stress evolution of different locations was investigated during the entire mining process. The maximum and minimum principal stresses at the same test point may have quite different changing trends. The maximum principle stress changes much more complex than the minimum principle stress, which may cause the permeability at the same location different along different directions. (3) The mining stope was divided into three concentric annuluses including the stress-decreased circle, stress increased circle and the original zone. The gas flow and enrichment characters were analyzed in different circles according to the stress value and direction, based on which, a gas control technology was developed by the combination of the boreholes and ventilation system. (4) It should be noted that there is no interface or joint element between the different layers (roof, coal seam and floor); the rock mass is modeled as an anisotropic media with three layers that have been completely welded together.
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Fig. 7. Maximum principal stress distribution and division of the mining stope (a) stress value in strike profile, (b) stress direction in strike profile, (c) stress value in dip profile, and (d) stress direction in dip profile.
Fig. 8. Comprehensive gas control technology according to gas flow and enrichment character.
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Fig. 9. Y-shaped ventilation system.
Acknowledgments This work was supported by the Fundamental Research Funds for the Central Universities (2014QNA02), Program for Changjiang Scholars and Innovative Research Team in University (Grant no. IRT13098), China Postdoctoral Science Foundation funded Project (2014M551057), National Natural Science Foundation of China (51404261), and Natural Science Foundation of Jiangsu province (BK20140196). References [1] Yang W, Lin B, Qu Y, Zhao S, Zhai C, Jia L, et al. Mechanism of strata deformation under protective seam and its application for relieved methane control. Int J Coal Geol 2011;85:300–6. [2] Li J, Su X, Na M, Yang H, Li J, Yu Y, et al. Influence of gas flow on thermal field and stress during growth of sapphire single crystal using the Kyropoulos method. Rare Metals 2006;25:260–6. [3] Wang JG, Kabir A, Liu JS, Chen ZW. Effects of non-Darcy flow on the performance of coal seam gas wells. Int J Coal Geol 2012;93:62–74. [4] Richards MJ. Longwall front abutment stress effects firedamp release. Min Sci Tech 1984;1:215–29. [5] Yang W, Lin B, Qu Y, Li Z, Zhai C, Jia L, et al. Stress evolution with time and space during mining of a coal seam. Int J Rock Mech Min Sci 2011;48:1145–52. [6] Hosseini N, Oraee K, Shahriar K, Goshtasbi K. Studying the stress redistribution around the longwall mining panel using passive seismic velocity tomography and geostatistical estimation. Arab J Geosci 2013;6:1407–16. [7] Luxbacher K, Westman E, Swanson P, Karfakis M. Three-dimensional timelapse velocity tomography of an underground longwall panel. Int J Rock Mech Min Sci 2008;45:478–85. [8] Mark C, Gale W, Oyler D, Chen J. Case history of the response of a longwall entry subjected to concentrated horizontal stress. Int J Rock Mech Min Sci 2007;44:210–21. [9] Suchowerska AM, Merifield RS, Carter JP. Vertical stress changes in multi-seam mining under supercritical longwall panels. Int J Rock Mech Min Sci 2013;61:306–20. [10] Islam MR, Shinjo R. Numerical simulation of stress distributions and displacements around an entry roadway with igneous intrusion and potential sources of seam gas emission of the Barapukuria coal mine, NW Bangladesh. Int J Coal Geol 2009;78:249–62. [11] Zhou H, Liu J, Xue D, Yi H, Xue J. Numerical simulation of gas flow process in mining-induced crack network. Int J Min Sci Tech 2012;22:793–9. [12] Whittles DN, Lowndes IS, Kingman SW, Yates C, Jobling S. Influence of geotechnical factors on gas flow experienced in a UK longwall coal mine panel. Int J Rock Mech Min Sci 2006;43:369–87.
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