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Field study on the load transfer mechanics associated with longwall coal retreat mining
International Journal of Rock Mechanics & Mining Sciences 124 (2019) 104141
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International Journal of Rock Mechanics and Mining Sciences journal homepage: http://www.elsevier.com/locate/ijrmms
Field study on the load transfer mechanics associated with longwall coal retreat mining Hongpu Kang a, b, *, Le Wu a, b, Fuqiang Gao a, b, Huawen Lv a, b, Jianzhong Li a, b a b
Mining & Designing Branch, China Coal Research Institute, Beijing, China State Key Laboratory of Coal Mining and Clean Utilization (China Coal Research Institute), Beijing, China
A R T I C L E I N F O
A B S T R A C T
Keywords: Longwall coal mine Load transfer mechanics Mining-induced stress In-situ monitoring
The two-entry longwall panel system is frequently used in underground coal mines in China. One of these entries serves only the current longwall panel while the other serves both current and subsequent longwall panels. The latter entry is built to survive the continuously high mining-induced stress generated by mining the current panel throughout the longwall mining process. The failure of the entry is controlled by the load transfer mechanics associated with the advancement of the longwall face. This paper presents a field study of the response of the entry during driving and longwall retreat phases. Extensive field data including entry convergence and changes in principal stresses are obtained for the Pingshu coal mine, China. A careful analysis of these data shows that the influence zone of the front abutment load on the trail entry can reach as far as 120 m outby the longwall face during panel mining. Bed separation can occur in competent fine-grained sandstone, which is 7.4 m deep into the trail entry roof. It is also found that the periodic collapse of the main roof above the mined-out area can continue when the longwall face has advanced 120 m away.
1. Introduction The two-entry longwall panel system is frequently used in under ground coal mines in China. This is particularly the case, for highly gassy coal seams, where sufficient fresh air is sent to the longwall face through the entries to meet minimum requirements set out by law, or for bumpprone areas where safety hazards should be avoided. One of the two entries only serves current longwall panel while the other serves both the current and next longwall panels. The latter entry is built to survive the continuously high mining-induced stresses generated by mining the current panel throughout the longwall mining process. Failure of the surrounding rock mass is generally inevitable. Challenges including vi olent face bursting and excessive tailgate convergence outby the face can be crippling to the ventilation. Understanding the performance of the entry is important to ground control design. The failure of an entry is controlled by the load transfer mechanics associated with the advancement of the longwall face. As the longwall face advances, the extraction of the coal alters the equilibrium of the insitu stress field, leading to abutment pressures being generated around the edge of the gob. The abutment pressure generated in front of the working face is termed the front abutment pressure whereas that
generated along both sides of the panel is called the side abutment pressure. The front abutment pressure increases sharply with distance into the un-mined coal, reaching a maximum value at a distance, and then gradually decreases to the pre-mining stress. The side abutment pressure has a similar trend with the distance from the gob edge. The magnitude and location of the maximum abutment pressure are related to the geological and mining conditions, including the depth and thickness of the coal seam, in situ stress conditions, lithology of the roof, longwall panel length, width, and mining speed. Many efforts have been made to study the nature and behavior of longwall mining-induced abutment loads. Peng1 suggested that the front abutment pressure in solid coal can first be detected at a distance of approximately 152 m outby the longwall face, begins to increase sharply at a distance of approximately 30 m, and then reaches a maximum value (i.e., maximum front abutment pressure) when the face is 0.9–6.0 m away. The maximum front abutment pressure ranges from 0.2 to 6.4 σ o (σ o ¼ γh, where h is the mining depth and γ is the weighted average unit weight of the overburden strata). Whittaker and Singh2 suggested that the maximum abutment pressure is in the range of 4–5 times σo for condi tions encountered in the United Kingdom. Majdi3 proposed an equation with which to calculate the coefficient of stress concentration due to
* Corresponding author. Mining & Designing Branch, China Coal Research Institute, Beijing, China. E-mail address: [email protected] (H. Kang). https://doi.org/10.1016/j.ijrmms.2019.104141 Received 22 March 2019; Received in revised form 18 October 2019; Accepted 23 October 2019 Available online 28 October 2019 1365-1609/© 2019 Elsevier Ltd. All rights reserved.
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
Fig. 1. Plan view of the study site showing the layout of the entries and panel as well as the locations of the monitoring sites for different purposes. Red and blue lines respectively refer to the part of the entry supported by patterns 1 and 2. A number in brackets indicates the distance from the monitoring site to the cut face. Not to scale. (For interpretation of the references to colour in this figure legend, the reader is referred to the Web version of this article.)
longwall mining: � � � �0:4 � Ep Mc ¼ 0:08h0:55 þ 0:70 0:002 Wo þ1 s Eg
rapidly within 10 m in front of the longwall face, and the maximum abutment pressure is approximately 2.0σ o when the coal seam is mined for 20–30 m. Through finite element modeling, Khanal et al.6 found that the maximum abutment pressure can exceed 4.0σo . Gao et al.7 per formed discrete element modeling based on a longwall coal panel case study in the German Ruhr mining district and found that the maximum abutment pressure reached 2.3σo . The above studies greatly improved our understanding of the local transfer mechanics associated with longwall coal mining. However, the influence zones and magnitude of the mining-induced stress are still not well defined or fully understood owing to the complexity of geological
(1)
where hs is the extracted coal seam thickness, Wo is the panel width, and Ep , and Eg are respectively the elastic moduli of the pillar and gob. Sheorey4 stated that the maximum pressure on the adjacent rib side and ahead of the face can reach 6.0σo and reduces to the pre-mining stress at a distance of 0.12H (where H is the depth of the coal seam). Yang et al.5 performed a field study and found that the abutment pressure changes
Fig. 2. Illustration of support patterns used in the trail entry. (a) First pattern. Roof support (left), rib support at the pillar side (right). (b) Second pattern. Roof support (left), rib support at the pillar side (right). For both patterns, the rib support at the solid coal side was the same as that at the pillar side except no cable bolt was installed. 2
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
2. Site description and instrumentation profile The study site is located at the Pingshu coal mine operating in the Yangquan coal field, China. A plan view of the longwall panel, showing the layout of the panel and entries, as well as the locations of monitoring stations is presented in Fig. 1. The coal seam had an average thickness of 2.3 m and a depth between 368 and 502 m, was sub-horizontal with a dip angle between 2� and 10� . The coal seam was extracted by longwall mining method. The width and length of longwall panel 81,115 were 180 and 1258 m, respectively. The longwall panel was accessed by a two-entry system with 5.0-m wide and 3.15-m high entries and cross cuts. The trail entry was driven parallel with the tailgate. The pillar between the two entries was 20-m wide. The crosscuts were in general angled at 90� , resulting in the formation of rectangular shaped chain pillars. The average chain pillar dimensions were 50 m � 20 m. The trail entry was supported by two different supporting patterns, see Fig. 2. The first pattern involved 1264-m driving from the trunk roadway and the second pattern involved 200-m driving from the cut face, see Fig. 1. For the first pattern, the roof support consisted of four 4.2-m long, 21.6-mm-in-diameter cable bolts and two 6.2-m long, 21.6mm-in-diameter cable bolts installed in a row with a steel strap. The cable bolts had spacing of at 0.9 m � 0.9 m along and across the entry. The rib support at the pillar side consisted of three 2.4-m long, 20-mmin-diameter round steel bolts and one 4.2-m long, 21.6-mm-in-diameter cable bolt. The rib support at the solid coal side consisted of four 2.4-m long, 20-mm-in-diameter round steel bolts and no cable bolt. For the second supporting pattern, the roof support consisted of five 2.4-m long, 22-mm-in-diameter resin anchored rebar. The rock bolts had spacing of 0.9 m � 0.9 m spacing along and across the entry, with W-shaped steel strap and welded screen. Cable bolts with a diameter of 21.6 mm and a length of 5.2 m were also installed in a two-three-two pattern to provide supplemental support. The spacing between cable rows was 1.8 m. The rib support at the pillar side consisted of four 2.4-m long, 22-mm-indiameter resin anchored rebars in a row. One cable bolt with a diam eter of 21.6 mm and a length of 4.2 m was also installed with a row spacing of 1.8 m. The rib support at the solid coal side consisted of four 2.4-m long, 22-mm-in-diameter resin anchored rebars in a row without any cable. In-situ stress measurements were conducted in the study entry adopting hydraulic fracturing. The measured in-situ maximum and minimum horizontal stresses were 11.46 and 6.25 MPa, respectively. The calculated vertical stress was 12.13 MPa. The in-situ maximum horizontal stress was oriented at about 90� to the longitudinal axis of the
Fig. 3. Representative lithology at the study site with the comprehensive strength obtained in borehole strength tests. (a) foof, (b) rib coal.
conditions and the longwall mining-induced overburden strata move ment. This paper presents a case study on the behavior of a longwall entry by means of field monitoring including consideration of entry deformation and stress changes. The objective of the study is to develop a better understanding of the loading mechanics associated with the advancement of the longwall face so that optimum support techniques can be applied and more effective support systems can be designed to reduce the risk of entry failure. The level of instrumentation used in this study is extensive and provides rare information to achieve this research objective.
Fig. 4. Locations of stress meters used for measuring changes in principal stresses as the longwall face advanced. The right plot is a section view of the left plot. Meters #1, #2, and #4 were located on the same cross section while meters #3 and #5 were located on the cross section ahead. 3
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
Fig. 5. Entry convergence monitored during the longwall retreat plotted against the distance of the longwall face. Positive numbers indicate the monitoring station is outby the longwall face, and negative numbers indicate the monitoring station is inby the longwall face. (a) Station 1 where the distance between the station and the cut face was 88 m, and (b) Station 2, where the distance between the station and the cut face was 223 m.
longwall panel, indicating that the entries were in the worst stress condition. Fig. 3 shows the representative lithology at the study site which was determined by borehole televiewer imaging performed in a 65-mm diameter borehole drilled vertically in the roof of the trial entry. Bore hole strength tests were carried out in the roof borehole and a horizontal borehole drilled in the coal rib to obtain the Coal Measure strength using in situ strength test equipment.8 A rod with a probe attached to the top was inserted into the borehole. The probe was connected to a high pressure pump with a pipe. Under the hydraulic pressure generated by the pump, the probe was pushed against the borehole wall until failure, identified by a sudden increase in the displacement of the probe and a constant pressure value. The pressure was recorded and used to estimate the rock strength, according to the relationship between the pressure and unconfined compressive strength (UCS) of the rock, which was established in a series of laboratory tests on different types of rock. As shown in Fig. 3 The immediate roof comprised sandy mudstone overlain
by medium-grained sandstone. Above was fine-grained sandstone overlain by coarse-grained sandstone. To capture the ground response of the trail entry during the longwall panel retreat phases, two convergence meters were installed at two sites at different distances from the cut face to measure deformation of the trail entry including roof sag, floor heave, and deformation at the two ribs. The distances from the cut face to sites 1 and 2 were 88 and 230 m, respectively. Site 1 was located in part of the trial entry that was sup ported by the first support pattern as shown in Fig. 2a, and site 2 was located in the part supported by the second support pattern as shown in Fig. 2b. In addition, five stress meters were installed at different locations of the main roof to measure changes in principal stresses as the longwall face approached and passed the monitoring station. Fig. 4 shows the locations of the stress meters. The site of the cells was 223 m from the cut face. Many instruments are available for measuring stress changes; e.g., deformation gages, strain cells, stiff cylindrical inclusions, solid and
Fig. 6. Field photographs showing severe floor heave caused by extensive abutment pressures. 4
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
Fig. 7. Field photographs showing severe rib deformation caused by extensive abutment pressures.
hollow deformable inclusions, and borehole stress gauges.9 Hollow deformable inclusions can measure both the magnitude and change of three-dimensional in-situ stresses.10 KX-81 hollow deformable in clusions were thus used in the present study to measure changes in the three principal stress components. Stress changes were obtained by converting the measured strain changes according to the deformation modulus of the rock within which the stress cells were installed. Rock cores were collected when borehole drilling was performed to install the stress meters. Standard rock sam ples with a diameter of 50 mm and a length of 100 mm were prepared from the rock cores and tested using servo loading equipment to mea sure the deformation modulus. The measured Young’s moduli of the fine-grained sandstone and coarse-grained sandstone were 14.79 and 3.48 GPa, respectively. The Poisson’s ratios of the two rocks were 0.24 and 0.41, respectively.
of mining-induced stresses as monitored by the convergence meters. It shows that the effects of the front abutment pressure began to manifest with increased entry deformation when the face was approximately 80 m away. The increase in the entry deformation remained gradual until the face was 20 m from the monitoring site, at which time the entry deformation increased significantly. After the longwall face passed the monitoring site, the trail entry deformed sharply. This sharp increase in the deformation continued when the longwall face was even 150 m outby the monitoring station. It can also be seen from Fig. 5 that the entry deformation was significantly asymmetrical. The floor heave was much larger than the roof sag and the deformation of the rib at the pillar side was larger than that at the solid coal side for both monitoring sites. The floor of the trial entry was severely damaged when the longwall face was around 200 m outby the monitoring site. Field observations show that when the highest floor heave reached about 0.9 m, the entry height reduced from its original 3.2 m–1.9 m, see Fig. 6. A 2-m long, 0.1-m wide, and 0.4-m deep tensile fracture formed as a result of this severe floor heave. Distinct damage was also observed in the rib at the pillar side, see Fig. 7. Even though the fractured coal was contained by the steel mesh, the inward deformation of the rib was so severe that the steel strap installed
3. Support and ground response during longwall retreat phases 3.1. Trial entry convergence Fig. 5 presents the deformation of the trail entry under the influence
Fig. 8. Changes in principal stress magnitudes (left) and directions (right) at Cell 1 located at the roof above the pillar center. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, while negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance. 5
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
Fig. 9. Changes in principal stress magnitudes (left) and directions (right) at Cell 2 located at the roof of the pillar. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance.
in the roof with the cable bolt bent. The rate and magnitude of the entry convergence at site 2 were remarkably greater than those at site 1. The roof sag, rib deformation at the pillar side, rib deformation at the solid coal side, and floor heave were respectively 71%, 68%, 64%, and 65% lower. One reason for the differences might be the different support patterns that were applied at the two sites (see Figs. 1 and 2), in that the first support pattern was more effective than the second pattern. Additionally, geological condi tions played an important role in the different deformations at the two sites. Site 1 was much closer to the cut face (i.e., a distance of 88 m) and fracturing of overburden rocks was not fully developed owing to the close boundary conditions.1 In other words, the development of frac tured zone was constrained by the unmined-out area at the cut face, and the mining-induced pressure applied to the entry at site 1 was therefore lower than that at site 1 where the fractured zone was fully developed because of the long distance (i.e. 230 m) to the cut face.
face continued to advance, a dramatic increase in all three stress com ponents was observed. At the end of monitoring, the increments of σ1 σ2 , and σ 3 were 34.1, 19.2, and 15.8 MPa, respectively, which were 2.8, 1.6, and 1.3 times the pre-mining vertical stress. The directions of the three stress components continuously changed as the longwall face approached and passed the cell location. The maximum principal stress change Δσ1 , initially sub-horizontal, increased in dip angle 50� in a south-westward direction as the long wall face approached. The dip of the minimum principal stress change Δσ3 decreased continuously from 50� to 15� in an eastward direction. The intermediate principal stress change Δσ2 first became dipper from 20� -25� to 80� –85� in a north-westward direction. As the longwall face approached and passed the cell location, Δσ2 rotated to the north-east and then rotated back to the north with the dip decreasing to 30� –40� . As the longwall face approached and passed the cell location, the vertical stress increased continuously, and the horizontal stress that was at a greater angle with the axis direction of the roadway initially decreased and then increased. The monitoring data suggest that the roof above the pillar remained stable. Fig. 9 presents the changes in principal stress magnitudes and di rections at Cell 2 located at the roof of the coal pillar. The horizontal distance from Cell 2 to the entry was 5.2 m, comparing with a distance of 9.5 m for Cell 1. After the installation of the cell, a distinct decrease was observed in all three stress components until the longwall face advanced 17 m, indicating a general deterioration of the roof that was driven by
3.2. Changes in principal stresses Fig. 8 presents changes in the principal stress magnitudes and di rections at Cell 1 which was located at the roof above the center of the coal pillar. The vector orientations are plotted on the lower hemisphere. A general subtle decrease in the magnitudes of the three stress compo nents was observed as the longwall face approached and passed by the location of Cell 1 until the longwall face was 30 m outby. As the longwall
Fig. 10. Changes in principal stress magnitudes (left) and directions (right) at Cell 3 located at the main roof of the pillar. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance. 6
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
Fig. 11. Periodic roof weighting shown as yield leg pressure monitored during the mining of 81115 panel. The leg pressure was obtained from the yield located in the middle of the longwall face, and the effect of side pillars as boundary conditions could that be ignored.
the entry excavation. It is believed that the effect of longwall mining was not involved in the roof deterioration because the longwall face was more than 200 m away and no further decrease in the stress components was observed as the longwall face continued to advance. The influence initiated when the longwall face was around 120 m inby the cell loca tion. Beyond this point, Δσ 3 gradually decreased and Δσ 1 gradually increased as the longwall face approached and passed the cell location. As the longwall face advanced, changes in the dip direction of the three stress change components were insignificant. Δσ2 remained in the eastwest direction with a shallow dip angle of 10� –20� , and Δσ 3 remained in the north-south direction with a shallow dip angle of 5� –20� . Δσ1 generally rotated continuously with a consistent sub-vertical direc tion. A dramatic decrease in σ 3 and σ 1 began when the longwall face was 70 m outby the cell location. The total drop of σ 3 was even greater than the maximum horizontal stress. This suggests that the coal pillar below Cell 2 had failed and the roof had ruptured in tension. Fig. 10 presents changes in principal stress magnitudes and
directions at Cell 3 located at the main roof fine-grained sandstone of the trail entry. No distinct changes in the three stress components were observed until the longwall face was 120 m away. As the longwall face continued to advance, a dramatic increase in σ 1 was observed. The maximum increase in σ 1 of 26 MPa occurred when the longwall face was 40 m outby the cell location. The change in σ2 was much lower than that of σ 1 during this retreat phase. A dramatic decrease in the minimum principal stress σ3 began when the longwall face was 120 m inby the cell location and continued to advance. The maximum decrease inσ3 was more than 30 MPa, leading to unreasonable tension. This can be explained by the non-linearity of the rock during this longwall retreat phase. It is interesting to note that a distinct periodic change was observed in all three stress components, which can be attributed to the periodic weighting of the main roof. The periodic roof weighting in terval was around 40–50 m, which agrees well with that observed from the shield leg pressure as shown in Fig. 11. The directions of the three stress change components continuously
Fig. 12. Changes in principal stress magnitudes (left) and directions (right) at Cell 5 located at the main roof coarse-grained sandstone of the pillar. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance. 7
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
a decrease in dip to 10� –15� . The dotted line indicates the start of the longwall face advance. Fig. 12 shows changes in principal stress magnitudes and directions at Cell 5 located at the main roof coarse-grained sandstone of the trail entry. Cell 5 was 5.6 m deeper than cell 3 into the main roof of the trial entry. Δσ 1 was essentially unchanged as the face approached and passed until the face was 40 m outby the cell location. An increase of 2.5 MPa was then observed when the longwall face continued to advance. The dip of Δσ 1 gradually increased from 10 to 30� during this longwall retreat phase. The magnitude of Δσ 3 was observed to decrease as the longwall face approached and passed the cell location. Δσ3 initially oriented in a south-westward direction with a deep dip of 60� –70� . As the longwall face approached, it suddenly dipped north-west with a deeper dip of 70–75� . As the longwall face passed, it dipped to the north with a much shallower angle of approximately 10� . A similar periodic change in the stress components was also observed as the longwall face reached 70 and 140 m outby the cell location. A direct comparison of the stress changes between Cells 3 and 5 located at different depths into the roof above the entry is shown in Fig. 13. It is clear that changes in all three stress components monitored by Cell 3 were greater than those monitored by Cell 5. This suggests that bed separation occurred within the entry roof at shallow depth (i.e. 7.4 m) under the influence of mining-induced stresses, while the deeper roof remained competent and sustained high abutment pressures. Cell 4 was located at the main roof coarse-grained sandstone above the solid coal side. The horizontal distances from Cell 4 to the 81115 trail entry and the 81115 ventilation entry were 9.5 and 34.5 m, respectively. Changes in principal stress magnitudes and directions at Cell 4 are shown in Fig. 14. In contrast to the cases of the other four cells, changes in the stress magnitudes and directions at Cell 4 were not drastic. This was due to Cell 4 being further from the mined-out area so that the mining-induced side abutment was not significant. To further evaluate the influence of the distance on the mininginduced side abutment pressure, a direct comparison of the stress changes of the three stress components is plotted in Fig. 15. The following conclusions can be drawn. Changes in σ1 at the roof above the coal pillar center (Cell 1) and the 81115 trail entry (Cell 3) were much greater than those at the roof above the solid coal (Cell 4) and the pillar side. At the location of Cell 1, stress changes mainly occurred in the sub-vertical direction after the longwall face had passed. At the location of Cell 3, stress changes mainly occurred in the sub-horizontal direction before the longwall face had passed. For each cell, Δσ2 was in the sub-horizontal direction. The greatest changes in σ2 occurred at Cell 1 and the majority of the changes happened after the longwall face had passed the cell location. At Cells 2, 3, and 4, Δσ 2 showed insignificant changes and tended to decrease after the longwall face had passed. At Cells 1 and 2, Δσ 3 was in the sub-horizontal direction. At Cell 3, Δσ3 was in the sub-vertical direction. At Cell 4, Δσ3 had a dip angle of 45� . The greatest increase in Δσ3 occurred at Cell 1 and the greatest decrease in Δσ 3 occurred at Cell 3. Fig. 13. Comparison of stress changes in Cells 3 and 5 at different depths into the entry roof. (a) Δσ 1 , (b) Δσ2 , and (c) Δσ 3 . On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face.
4. Discussion The extraction of a coal seam and the subsequent roof caving alters the in-situ stress field, resulting in stress concentration around openings. Mining-induced stress is the cause of many geological problems including severe entry convergence,11,12 roof collapse13 and coal bursts.14–16 Understanding the evolution of mining-induced stress and affecting factors is therefore important in solving such geological problems. The range of mining-induced stress (i.e., the distance to the longwall face from where mining-induced stress is first detected) is related to many factors, including geological and geotechnical condi tions. It has been reported that in western United States coal mines under a cover of 402 m, abutment pressure can be detected when the longwall face is 107–198 m away17 and. In the Czech part of the Upper
changed as the longwall face approached and passed the cell location. Δσ1 , initially horizontal, dipped to 60� as the longwall face approached. As the longwall face passed, it gradually rotated back to a sub-vertical direction. Δσ 3 first became shallower in a northward direction and then rotated back with a deeper dip of 70� –75� . As the longwall face approached and passed the cell location, it rotated from north to west. Δσ2 first became shallower from 50� -65� to 15� –20� in a southward di rection. As the longwall face approached and passed the cell location, Δσ2 rotated to east with an initial decrease in dip to 40� –50� followed by 8
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
Fig. 14. Changes in principal stress magnitudes (left) and directions (right) at Cell 4 located at the main roof coarse-grained sandstone of the pillar. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance.
Fig. 15. Influence of distance on mining-induced side abutment pressure. (a) Δσ 1 , (b) Δσ 2 , and (c) Δσ 3 . The curve numbers indicate the distance from the longwall face to the site location of the stress meters. Positive values indicate the longwall face was outby the site location, and negative values indicate the longwall face was inby the site location.
was 100 m away.18 At a mining depth of 150 m in Emerald mine, front abutment pressure was observed when the longwall face was 33 m away.19 In the present study, the depth of the coal seam was 368–502 m, and there was front abutment pressure when the longwall face was 120 m away. Despite the scatter of data that may be due to the local
Silesian Coal Basin at a mining depth of 740 m, an obvious increase in the front abutment pressure was observed when the longwall face was 200 m away and a larger increase was observed when the longwall face was 40 m away. In the Tashan coal mine, China with a overburden depth of 300–500 m, abutment pressure was observed when the longwall face 9
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geological and geotechnical conditions, it is found that the mining depth is a key factor controlling the range of mining-induced stress. The greater the depth, the greater the range of the mining-induced stress. After a coal seam is mined out, the caving of the immediate roof and fracturing of the overburden rocks is a dynamic process that occurs as the longwall face continues to advance, resulting in a gradual change in the mining-induced stress applied to the unmined-out area, i.e., the pillar and entry that will serve the next longwall panel. The monitoring of changes in principal stresses in the present study showed distinct stress changes after the longwall face passed (see Figs. 8–10, 12, and 14), these were a direct indication of the gradual fracturing of the over burden rocks. This finding was similar to a result reported in the liter ature18 that the field monitored pillar stress continued to increase as the longwall face passed by a monitoring station and reached a peak value when the longwall face was 360–379 m away. The continuous advance of the longwall panel seemed to play a role in the gradual fracturing of overburden rocks, as suggested by the distinct periodic change in the principal stresses observed by all five cells. This periodic characteristic of the overburden rock fracturing above the back of the gob plays an important role in the stress condition of the next longwall panel and the surface subsidence resulting from longwall mining.20 Guo et al.21 found that fracturing can reach all the way to the surface when the longwall width-to-depth ratio is greater than 0.75. In the present study, the width of the longwall panel was 180 m, and the depth was 368–502 m, giving a longwall width-to-depth ratio of 0.36–0.49. This suggests that fracturing of the overburden does not reach the ground surface if Guo’s result applies to the present case. The fracturing pattern and process of the overburden varies depending on the location about the panel, leading to a non-uniform distribution of the fractured zone and abutment pressure across the panel width.1 It is clear from Fig. 15 that the abutment pressure is substantially different at different locations from the longwall panel. The greatest increase in the maximum principal stress occurred at Cell 1, which was closest to the panel, suggesting that the peak side abutment pressure occurred in the periphery of the gob with a range of around 11 m. This finding provides fundamental data for the implementation of stress relief measures including blasting and hydraulic fracturing.22,23
maximize the efficiency of a multiple-entry longwall system in terms of ventilation, ground control, economic, and safety considerations. Acknowledgments This work has been supported by the National Key Research and Development Program of China (grant no. 2017YFC0603003). Appendix A. Supplementary data Supplementary data to this article can be found online at https://doi. org/10.1016/j.ijrmms.2019.104141. References 1 Peng S. Coal Mine Ground Control. 2rd ed. New York: Wiley; 1986. second ed. 2 Whittaker BN, Singh RN. Evaluation of the design requirements and performance of gate roadways. Min Eng. 1979;138:535–548. 3 Majdi A. The Stability of Face Access Tunnel in the Deep Soft Rocks of Coal Mining. 1988. 4 Sheorey PR. Design of coal pillar arrays and chain pillars. In: Fairhurst C, ed. Analysis and Design Methods. Oxford: Pergamon; 1993:631–670. https://doi.org/10.1016/ B978-0-08-040615-2.50030-7. 5 Yang W, Lin B, Qu Y, et al. Stress evolution with time and space during mining of a coal seam. Int J Rock Mech Min Sci. 2011;48(7):1145–1152. https://doi.org/ 10.1016/j.ijrmms.2011.07.006. 6 Khanal M, Adhikary D, Balusu R. Evaluation of mine scale longwall top coal caving parameters using continuum analysis. Min Sci Technol China. 2011;21(6):787–796. https://doi.org/10.1016/j.mstc.2011.06.027. 7 Gao F, Stead D, Coggan J. Evaluation of coal longwall caving characteristics using an innovative UDEC Trigon approach. Comput Geotech. 2014;55:448–460. https://doi. org/10.1016/j.compgeo.2013.09.020. 8 Gao F, Stead D, Kang H, Wu Y. Discrete element modelling of deformation and damage of a roadway driven along an unstable goaf — a case study. Int J Coal Geol. 2014;127:100–110. https://doi.org/10.1016/j.coal.2014.02.010. 9 Amadei B, Stephansson O. Rock Stress and its Measurement. London: Chapman & Hall; 1997. 10 Mills KW, Pender MJ. A soft inclusion instrument for in situ stress measurement in coal. In: International Society for Rock Mechanics and Rock Engineering; 1986. http s://www.onepetro.org/conference-paper/ISRM-IS-1986-025. Accessed March 4, 2019. 11 Kang HP, Lin J, Fan MJ. Investigation on support pattern of a coal mine roadway within soft rocks — a case study. Int J Coal Geol. 2015;140:31–40. https://doi.org/ 10.1016/j.coal.2015.01.003. 12 Peng SS. Ground Control Failures: A Pictorial View of Case Studies. West Virginia University; 2007. 13 Kang H, Lou J, Gao F, Yang J, Li J. A physical and numerical investigation of sudden massive roof collapse during longwall coal retreat mining. Int J Coal Geol. 2018;188: 25–36. https://doi.org/10.1016/j.coal.2018.01.013. 14 Brauner G. Rockbursts in Coal Mines and Their Prevention. Rotterdam: A.A. Belkema; 1994. https://www.osti.gov/etdeweb/biblio/143105. Accessed October 9, 2019. 15 Konicek P, Waclawik P. Stress changes and seismicity monitoring of hard coal longwall mining in high rockburst risk areas. Tunn Undergr Space Technol. 2018;81: 237–251. https://doi.org/10.1016/j.tust.2018.07.019. 16 Pt� a�cek J. Rockburst in ostrava-karvina coalfield. Procedia Eng. 2017;191:1144–1151. https://doi.org/10.1016/j.proeng.2017.05.289. 17 Chen J, Mishra M, Zahl E, Dunford J, Thomas R. Longwall mining-induced abutment loads and their impacts on pillar design and entry stability. In: Proceedings of the 21st International Conference on Ground Control in Mining. WV, USA: Morgantown; 2002: 11–17. 18 Yu B, Zhang Z, Kuang T, Liu J. Stress changes and deformation monitoring of longwall coal pillars located in weak ground. Rock Mech Rock Eng. 2016;49(8): 3293–3305. https://doi.org/10.1007/s00603-016-0970-8. 19 Barczak TM, Tadolini SC, Zhang P. Evaluation of support and ground response as longwall face advances into and widens pre-driven recovery room. In: Proceedings of the 26th International Conference on Ground Control in Mining. WV: Morgantown; 2007:160–172. 20 Jeran PW, Trevits MA. Timing and Duration of Subsidence Due to Longwall Mining. U.S. Department of the Interior; 1995. https://www.cdc.gov/niosh/mining/works/cov ersheet1746.html. Accessed October 11, 2019. 21 Guo H, Adhikary DP, Gabeva D. Hydrogelogical Response to Longwall Mining. 2007: 148. 22 Gale WJ, Nemcik J, Upfold RW. Application of stress control methods to underground coal mine design in high lateral stress fields. In: Proc 6th International Conference on Rock Mechanics. vol. 25. 1987:897–900. https://doi.org/10.1016/ 0148-9062(88)91803-7. Montreal. 23 Kang H, Lv H, Gao F, Meng X, Feng Y. Understanding mechanisms of destressing mining-induced stresses using hydraulic fracturing. Int J Coal Geol. 2018;196:19–28. https://doi.org/10.1016/j.coal.2018.06.023.
5. Conclusions The following conclusions are drawn from the field monitoring data, which provide valuable information to better understand the load transfer mechanics associated with longwall retreat mining. Mining depth is a key factor controlling the range of mining-induced stress. The greater the depth, the greater the range of the mininginduced stress. The influence zone of the front abutment load on the trail entry reached as far as 120 m outby the longwall face during panel mining. The influence of the front abutment pressure, however, was not significant until the longwall face was 20 m inby. As the longwall face continued to approach and pass, there was a dramatic increase in the deformation of the trail entry. Owing to the high abutment pressure, bed separation could occur in the competent fine-grained sandstone that was 7.4 m deep into the trail entry roof. The first support pattern consisting of high-strength rock bolts and cable bolts installed with a W-shaped steel strap and welded screen was found to be effective in sustaining the integrity of the trail entry. Periodic collapse of the main roof above the mined-out area continued when the longwall face had advanced 120 m away. The peak side abutment pressure occurred in the periphery of the gob with a range of around 11 m. The results of the present study improve our understanding of the local transfer mechanics associated with longwall coal mining. Such understanding allows the development of design criteria that will help
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International Journal of Rock Mechanics and Mining Sciences journal homepage: http://www.elsevier.com/locate/ijrmms
Field study on the load transfer mechanics associated with longwall coal retreat mining Hongpu Kang a, b, *, Le Wu a, b, Fuqiang Gao a, b, Huawen Lv a, b, Jianzhong Li a, b a b
Mining & Designing Branch, China Coal Research Institute, Beijing, China State Key Laboratory of Coal Mining and Clean Utilization (China Coal Research Institute), Beijing, China
A R T I C L E I N F O
A B S T R A C T
Keywords: Longwall coal mine Load transfer mechanics Mining-induced stress In-situ monitoring
The two-entry longwall panel system is frequently used in underground coal mines in China. One of these entries serves only the current longwall panel while the other serves both current and subsequent longwall panels. The latter entry is built to survive the continuously high mining-induced stress generated by mining the current panel throughout the longwall mining process. The failure of the entry is controlled by the load transfer mechanics associated with the advancement of the longwall face. This paper presents a field study of the response of the entry during driving and longwall retreat phases. Extensive field data including entry convergence and changes in principal stresses are obtained for the Pingshu coal mine, China. A careful analysis of these data shows that the influence zone of the front abutment load on the trail entry can reach as far as 120 m outby the longwall face during panel mining. Bed separation can occur in competent fine-grained sandstone, which is 7.4 m deep into the trail entry roof. It is also found that the periodic collapse of the main roof above the mined-out area can continue when the longwall face has advanced 120 m away.
1. Introduction The two-entry longwall panel system is frequently used in under ground coal mines in China. This is particularly the case, for highly gassy coal seams, where sufficient fresh air is sent to the longwall face through the entries to meet minimum requirements set out by law, or for bumpprone areas where safety hazards should be avoided. One of the two entries only serves current longwall panel while the other serves both the current and next longwall panels. The latter entry is built to survive the continuously high mining-induced stresses generated by mining the current panel throughout the longwall mining process. Failure of the surrounding rock mass is generally inevitable. Challenges including vi olent face bursting and excessive tailgate convergence outby the face can be crippling to the ventilation. Understanding the performance of the entry is important to ground control design. The failure of an entry is controlled by the load transfer mechanics associated with the advancement of the longwall face. As the longwall face advances, the extraction of the coal alters the equilibrium of the insitu stress field, leading to abutment pressures being generated around the edge of the gob. The abutment pressure generated in front of the working face is termed the front abutment pressure whereas that
generated along both sides of the panel is called the side abutment pressure. The front abutment pressure increases sharply with distance into the un-mined coal, reaching a maximum value at a distance, and then gradually decreases to the pre-mining stress. The side abutment pressure has a similar trend with the distance from the gob edge. The magnitude and location of the maximum abutment pressure are related to the geological and mining conditions, including the depth and thickness of the coal seam, in situ stress conditions, lithology of the roof, longwall panel length, width, and mining speed. Many efforts have been made to study the nature and behavior of longwall mining-induced abutment loads. Peng1 suggested that the front abutment pressure in solid coal can first be detected at a distance of approximately 152 m outby the longwall face, begins to increase sharply at a distance of approximately 30 m, and then reaches a maximum value (i.e., maximum front abutment pressure) when the face is 0.9–6.0 m away. The maximum front abutment pressure ranges from 0.2 to 6.4 σ o (σ o ¼ γh, where h is the mining depth and γ is the weighted average unit weight of the overburden strata). Whittaker and Singh2 suggested that the maximum abutment pressure is in the range of 4–5 times σo for condi tions encountered in the United Kingdom. Majdi3 proposed an equation with which to calculate the coefficient of stress concentration due to
* Corresponding author. Mining & Designing Branch, China Coal Research Institute, Beijing, China. E-mail address: [email protected] (H. Kang). https://doi.org/10.1016/j.ijrmms.2019.104141 Received 22 March 2019; Received in revised form 18 October 2019; Accepted 23 October 2019 Available online 28 October 2019 1365-1609/© 2019 Elsevier Ltd. All rights reserved.
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Fig. 1. Plan view of the study site showing the layout of the entries and panel as well as the locations of the monitoring sites for different purposes. Red and blue lines respectively refer to the part of the entry supported by patterns 1 and 2. A number in brackets indicates the distance from the monitoring site to the cut face. Not to scale. (For interpretation of the references to colour in this figure legend, the reader is referred to the Web version of this article.)
longwall mining: � � � �0:4 � Ep Mc ¼ 0:08h0:55 þ 0:70 0:002 Wo þ1 s Eg
rapidly within 10 m in front of the longwall face, and the maximum abutment pressure is approximately 2.0σ o when the coal seam is mined for 20–30 m. Through finite element modeling, Khanal et al.6 found that the maximum abutment pressure can exceed 4.0σo . Gao et al.7 per formed discrete element modeling based on a longwall coal panel case study in the German Ruhr mining district and found that the maximum abutment pressure reached 2.3σo . The above studies greatly improved our understanding of the local transfer mechanics associated with longwall coal mining. However, the influence zones and magnitude of the mining-induced stress are still not well defined or fully understood owing to the complexity of geological
(1)
where hs is the extracted coal seam thickness, Wo is the panel width, and Ep , and Eg are respectively the elastic moduli of the pillar and gob. Sheorey4 stated that the maximum pressure on the adjacent rib side and ahead of the face can reach 6.0σo and reduces to the pre-mining stress at a distance of 0.12H (where H is the depth of the coal seam). Yang et al.5 performed a field study and found that the abutment pressure changes
Fig. 2. Illustration of support patterns used in the trail entry. (a) First pattern. Roof support (left), rib support at the pillar side (right). (b) Second pattern. Roof support (left), rib support at the pillar side (right). For both patterns, the rib support at the solid coal side was the same as that at the pillar side except no cable bolt was installed. 2
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2. Site description and instrumentation profile The study site is located at the Pingshu coal mine operating in the Yangquan coal field, China. A plan view of the longwall panel, showing the layout of the panel and entries, as well as the locations of monitoring stations is presented in Fig. 1. The coal seam had an average thickness of 2.3 m and a depth between 368 and 502 m, was sub-horizontal with a dip angle between 2� and 10� . The coal seam was extracted by longwall mining method. The width and length of longwall panel 81,115 were 180 and 1258 m, respectively. The longwall panel was accessed by a two-entry system with 5.0-m wide and 3.15-m high entries and cross cuts. The trail entry was driven parallel with the tailgate. The pillar between the two entries was 20-m wide. The crosscuts were in general angled at 90� , resulting in the formation of rectangular shaped chain pillars. The average chain pillar dimensions were 50 m � 20 m. The trail entry was supported by two different supporting patterns, see Fig. 2. The first pattern involved 1264-m driving from the trunk roadway and the second pattern involved 200-m driving from the cut face, see Fig. 1. For the first pattern, the roof support consisted of four 4.2-m long, 21.6-mm-in-diameter cable bolts and two 6.2-m long, 21.6mm-in-diameter cable bolts installed in a row with a steel strap. The cable bolts had spacing of at 0.9 m � 0.9 m along and across the entry. The rib support at the pillar side consisted of three 2.4-m long, 20-mmin-diameter round steel bolts and one 4.2-m long, 21.6-mm-in-diameter cable bolt. The rib support at the solid coal side consisted of four 2.4-m long, 20-mm-in-diameter round steel bolts and no cable bolt. For the second supporting pattern, the roof support consisted of five 2.4-m long, 22-mm-in-diameter resin anchored rebar. The rock bolts had spacing of 0.9 m � 0.9 m spacing along and across the entry, with W-shaped steel strap and welded screen. Cable bolts with a diameter of 21.6 mm and a length of 5.2 m were also installed in a two-three-two pattern to provide supplemental support. The spacing between cable rows was 1.8 m. The rib support at the pillar side consisted of four 2.4-m long, 22-mm-indiameter resin anchored rebars in a row. One cable bolt with a diam eter of 21.6 mm and a length of 4.2 m was also installed with a row spacing of 1.8 m. The rib support at the solid coal side consisted of four 2.4-m long, 22-mm-in-diameter resin anchored rebars in a row without any cable. In-situ stress measurements were conducted in the study entry adopting hydraulic fracturing. The measured in-situ maximum and minimum horizontal stresses were 11.46 and 6.25 MPa, respectively. The calculated vertical stress was 12.13 MPa. The in-situ maximum horizontal stress was oriented at about 90� to the longitudinal axis of the
Fig. 3. Representative lithology at the study site with the comprehensive strength obtained in borehole strength tests. (a) foof, (b) rib coal.
conditions and the longwall mining-induced overburden strata move ment. This paper presents a case study on the behavior of a longwall entry by means of field monitoring including consideration of entry deformation and stress changes. The objective of the study is to develop a better understanding of the loading mechanics associated with the advancement of the longwall face so that optimum support techniques can be applied and more effective support systems can be designed to reduce the risk of entry failure. The level of instrumentation used in this study is extensive and provides rare information to achieve this research objective.
Fig. 4. Locations of stress meters used for measuring changes in principal stresses as the longwall face advanced. The right plot is a section view of the left plot. Meters #1, #2, and #4 were located on the same cross section while meters #3 and #5 were located on the cross section ahead. 3
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Fig. 5. Entry convergence monitored during the longwall retreat plotted against the distance of the longwall face. Positive numbers indicate the monitoring station is outby the longwall face, and negative numbers indicate the monitoring station is inby the longwall face. (a) Station 1 where the distance between the station and the cut face was 88 m, and (b) Station 2, where the distance between the station and the cut face was 223 m.
longwall panel, indicating that the entries were in the worst stress condition. Fig. 3 shows the representative lithology at the study site which was determined by borehole televiewer imaging performed in a 65-mm diameter borehole drilled vertically in the roof of the trial entry. Bore hole strength tests were carried out in the roof borehole and a horizontal borehole drilled in the coal rib to obtain the Coal Measure strength using in situ strength test equipment.8 A rod with a probe attached to the top was inserted into the borehole. The probe was connected to a high pressure pump with a pipe. Under the hydraulic pressure generated by the pump, the probe was pushed against the borehole wall until failure, identified by a sudden increase in the displacement of the probe and a constant pressure value. The pressure was recorded and used to estimate the rock strength, according to the relationship between the pressure and unconfined compressive strength (UCS) of the rock, which was established in a series of laboratory tests on different types of rock. As shown in Fig. 3 The immediate roof comprised sandy mudstone overlain
by medium-grained sandstone. Above was fine-grained sandstone overlain by coarse-grained sandstone. To capture the ground response of the trail entry during the longwall panel retreat phases, two convergence meters were installed at two sites at different distances from the cut face to measure deformation of the trail entry including roof sag, floor heave, and deformation at the two ribs. The distances from the cut face to sites 1 and 2 were 88 and 230 m, respectively. Site 1 was located in part of the trial entry that was sup ported by the first support pattern as shown in Fig. 2a, and site 2 was located in the part supported by the second support pattern as shown in Fig. 2b. In addition, five stress meters were installed at different locations of the main roof to measure changes in principal stresses as the longwall face approached and passed the monitoring station. Fig. 4 shows the locations of the stress meters. The site of the cells was 223 m from the cut face. Many instruments are available for measuring stress changes; e.g., deformation gages, strain cells, stiff cylindrical inclusions, solid and
Fig. 6. Field photographs showing severe floor heave caused by extensive abutment pressures. 4
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Fig. 7. Field photographs showing severe rib deformation caused by extensive abutment pressures.
hollow deformable inclusions, and borehole stress gauges.9 Hollow deformable inclusions can measure both the magnitude and change of three-dimensional in-situ stresses.10 KX-81 hollow deformable in clusions were thus used in the present study to measure changes in the three principal stress components. Stress changes were obtained by converting the measured strain changes according to the deformation modulus of the rock within which the stress cells were installed. Rock cores were collected when borehole drilling was performed to install the stress meters. Standard rock sam ples with a diameter of 50 mm and a length of 100 mm were prepared from the rock cores and tested using servo loading equipment to mea sure the deformation modulus. The measured Young’s moduli of the fine-grained sandstone and coarse-grained sandstone were 14.79 and 3.48 GPa, respectively. The Poisson’s ratios of the two rocks were 0.24 and 0.41, respectively.
of mining-induced stresses as monitored by the convergence meters. It shows that the effects of the front abutment pressure began to manifest with increased entry deformation when the face was approximately 80 m away. The increase in the entry deformation remained gradual until the face was 20 m from the monitoring site, at which time the entry deformation increased significantly. After the longwall face passed the monitoring site, the trail entry deformed sharply. This sharp increase in the deformation continued when the longwall face was even 150 m outby the monitoring station. It can also be seen from Fig. 5 that the entry deformation was significantly asymmetrical. The floor heave was much larger than the roof sag and the deformation of the rib at the pillar side was larger than that at the solid coal side for both monitoring sites. The floor of the trial entry was severely damaged when the longwall face was around 200 m outby the monitoring site. Field observations show that when the highest floor heave reached about 0.9 m, the entry height reduced from its original 3.2 m–1.9 m, see Fig. 6. A 2-m long, 0.1-m wide, and 0.4-m deep tensile fracture formed as a result of this severe floor heave. Distinct damage was also observed in the rib at the pillar side, see Fig. 7. Even though the fractured coal was contained by the steel mesh, the inward deformation of the rib was so severe that the steel strap installed
3. Support and ground response during longwall retreat phases 3.1. Trial entry convergence Fig. 5 presents the deformation of the trail entry under the influence
Fig. 8. Changes in principal stress magnitudes (left) and directions (right) at Cell 1 located at the roof above the pillar center. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, while negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance. 5
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Fig. 9. Changes in principal stress magnitudes (left) and directions (right) at Cell 2 located at the roof of the pillar. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance.
in the roof with the cable bolt bent. The rate and magnitude of the entry convergence at site 2 were remarkably greater than those at site 1. The roof sag, rib deformation at the pillar side, rib deformation at the solid coal side, and floor heave were respectively 71%, 68%, 64%, and 65% lower. One reason for the differences might be the different support patterns that were applied at the two sites (see Figs. 1 and 2), in that the first support pattern was more effective than the second pattern. Additionally, geological condi tions played an important role in the different deformations at the two sites. Site 1 was much closer to the cut face (i.e., a distance of 88 m) and fracturing of overburden rocks was not fully developed owing to the close boundary conditions.1 In other words, the development of frac tured zone was constrained by the unmined-out area at the cut face, and the mining-induced pressure applied to the entry at site 1 was therefore lower than that at site 1 where the fractured zone was fully developed because of the long distance (i.e. 230 m) to the cut face.
face continued to advance, a dramatic increase in all three stress com ponents was observed. At the end of monitoring, the increments of σ1 σ2 , and σ 3 were 34.1, 19.2, and 15.8 MPa, respectively, which were 2.8, 1.6, and 1.3 times the pre-mining vertical stress. The directions of the three stress components continuously changed as the longwall face approached and passed the cell location. The maximum principal stress change Δσ1 , initially sub-horizontal, increased in dip angle 50� in a south-westward direction as the long wall face approached. The dip of the minimum principal stress change Δσ3 decreased continuously from 50� to 15� in an eastward direction. The intermediate principal stress change Δσ2 first became dipper from 20� -25� to 80� –85� in a north-westward direction. As the longwall face approached and passed the cell location, Δσ2 rotated to the north-east and then rotated back to the north with the dip decreasing to 30� –40� . As the longwall face approached and passed the cell location, the vertical stress increased continuously, and the horizontal stress that was at a greater angle with the axis direction of the roadway initially decreased and then increased. The monitoring data suggest that the roof above the pillar remained stable. Fig. 9 presents the changes in principal stress magnitudes and di rections at Cell 2 located at the roof of the coal pillar. The horizontal distance from Cell 2 to the entry was 5.2 m, comparing with a distance of 9.5 m for Cell 1. After the installation of the cell, a distinct decrease was observed in all three stress components until the longwall face advanced 17 m, indicating a general deterioration of the roof that was driven by
3.2. Changes in principal stresses Fig. 8 presents changes in the principal stress magnitudes and di rections at Cell 1 which was located at the roof above the center of the coal pillar. The vector orientations are plotted on the lower hemisphere. A general subtle decrease in the magnitudes of the three stress compo nents was observed as the longwall face approached and passed by the location of Cell 1 until the longwall face was 30 m outby. As the longwall
Fig. 10. Changes in principal stress magnitudes (left) and directions (right) at Cell 3 located at the main roof of the pillar. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance. 6
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Fig. 11. Periodic roof weighting shown as yield leg pressure monitored during the mining of 81115 panel. The leg pressure was obtained from the yield located in the middle of the longwall face, and the effect of side pillars as boundary conditions could that be ignored.
the entry excavation. It is believed that the effect of longwall mining was not involved in the roof deterioration because the longwall face was more than 200 m away and no further decrease in the stress components was observed as the longwall face continued to advance. The influence initiated when the longwall face was around 120 m inby the cell loca tion. Beyond this point, Δσ 3 gradually decreased and Δσ 1 gradually increased as the longwall face approached and passed the cell location. As the longwall face advanced, changes in the dip direction of the three stress change components were insignificant. Δσ2 remained in the eastwest direction with a shallow dip angle of 10� –20� , and Δσ 3 remained in the north-south direction with a shallow dip angle of 5� –20� . Δσ1 generally rotated continuously with a consistent sub-vertical direc tion. A dramatic decrease in σ 3 and σ 1 began when the longwall face was 70 m outby the cell location. The total drop of σ 3 was even greater than the maximum horizontal stress. This suggests that the coal pillar below Cell 2 had failed and the roof had ruptured in tension. Fig. 10 presents changes in principal stress magnitudes and
directions at Cell 3 located at the main roof fine-grained sandstone of the trail entry. No distinct changes in the three stress components were observed until the longwall face was 120 m away. As the longwall face continued to advance, a dramatic increase in σ 1 was observed. The maximum increase in σ 1 of 26 MPa occurred when the longwall face was 40 m outby the cell location. The change in σ2 was much lower than that of σ 1 during this retreat phase. A dramatic decrease in the minimum principal stress σ3 began when the longwall face was 120 m inby the cell location and continued to advance. The maximum decrease inσ3 was more than 30 MPa, leading to unreasonable tension. This can be explained by the non-linearity of the rock during this longwall retreat phase. It is interesting to note that a distinct periodic change was observed in all three stress components, which can be attributed to the periodic weighting of the main roof. The periodic roof weighting in terval was around 40–50 m, which agrees well with that observed from the shield leg pressure as shown in Fig. 11. The directions of the three stress change components continuously
Fig. 12. Changes in principal stress magnitudes (left) and directions (right) at Cell 5 located at the main roof coarse-grained sandstone of the pillar. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance. 7
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a decrease in dip to 10� –15� . The dotted line indicates the start of the longwall face advance. Fig. 12 shows changes in principal stress magnitudes and directions at Cell 5 located at the main roof coarse-grained sandstone of the trail entry. Cell 5 was 5.6 m deeper than cell 3 into the main roof of the trial entry. Δσ 1 was essentially unchanged as the face approached and passed until the face was 40 m outby the cell location. An increase of 2.5 MPa was then observed when the longwall face continued to advance. The dip of Δσ 1 gradually increased from 10 to 30� during this longwall retreat phase. The magnitude of Δσ 3 was observed to decrease as the longwall face approached and passed the cell location. Δσ3 initially oriented in a south-westward direction with a deep dip of 60� –70� . As the longwall face approached, it suddenly dipped north-west with a deeper dip of 70–75� . As the longwall face passed, it dipped to the north with a much shallower angle of approximately 10� . A similar periodic change in the stress components was also observed as the longwall face reached 70 and 140 m outby the cell location. A direct comparison of the stress changes between Cells 3 and 5 located at different depths into the roof above the entry is shown in Fig. 13. It is clear that changes in all three stress components monitored by Cell 3 were greater than those monitored by Cell 5. This suggests that bed separation occurred within the entry roof at shallow depth (i.e. 7.4 m) under the influence of mining-induced stresses, while the deeper roof remained competent and sustained high abutment pressures. Cell 4 was located at the main roof coarse-grained sandstone above the solid coal side. The horizontal distances from Cell 4 to the 81115 trail entry and the 81115 ventilation entry were 9.5 and 34.5 m, respectively. Changes in principal stress magnitudes and directions at Cell 4 are shown in Fig. 14. In contrast to the cases of the other four cells, changes in the stress magnitudes and directions at Cell 4 were not drastic. This was due to Cell 4 being further from the mined-out area so that the mining-induced side abutment was not significant. To further evaluate the influence of the distance on the mininginduced side abutment pressure, a direct comparison of the stress changes of the three stress components is plotted in Fig. 15. The following conclusions can be drawn. Changes in σ1 at the roof above the coal pillar center (Cell 1) and the 81115 trail entry (Cell 3) were much greater than those at the roof above the solid coal (Cell 4) and the pillar side. At the location of Cell 1, stress changes mainly occurred in the sub-vertical direction after the longwall face had passed. At the location of Cell 3, stress changes mainly occurred in the sub-horizontal direction before the longwall face had passed. For each cell, Δσ2 was in the sub-horizontal direction. The greatest changes in σ2 occurred at Cell 1 and the majority of the changes happened after the longwall face had passed the cell location. At Cells 2, 3, and 4, Δσ 2 showed insignificant changes and tended to decrease after the longwall face had passed. At Cells 1 and 2, Δσ 3 was in the sub-horizontal direction. At Cell 3, Δσ3 was in the sub-vertical direction. At Cell 4, Δσ3 had a dip angle of 45� . The greatest increase in Δσ3 occurred at Cell 1 and the greatest decrease in Δσ 3 occurred at Cell 3. Fig. 13. Comparison of stress changes in Cells 3 and 5 at different depths into the entry roof. (a) Δσ 1 , (b) Δσ2 , and (c) Δσ 3 . On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face.
4. Discussion The extraction of a coal seam and the subsequent roof caving alters the in-situ stress field, resulting in stress concentration around openings. Mining-induced stress is the cause of many geological problems including severe entry convergence,11,12 roof collapse13 and coal bursts.14–16 Understanding the evolution of mining-induced stress and affecting factors is therefore important in solving such geological problems. The range of mining-induced stress (i.e., the distance to the longwall face from where mining-induced stress is first detected) is related to many factors, including geological and geotechnical condi tions. It has been reported that in western United States coal mines under a cover of 402 m, abutment pressure can be detected when the longwall face is 107–198 m away17 and. In the Czech part of the Upper
changed as the longwall face approached and passed the cell location. Δσ1 , initially horizontal, dipped to 60� as the longwall face approached. As the longwall face passed, it gradually rotated back to a sub-vertical direction. Δσ 3 first became shallower in a northward direction and then rotated back with a deeper dip of 70� –75� . As the longwall face approached and passed the cell location, it rotated from north to west. Δσ2 first became shallower from 50� -65� to 15� –20� in a southward di rection. As the longwall face approached and passed the cell location, Δσ2 rotated to east with an initial decrease in dip to 40� –50� followed by 8
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
Fig. 14. Changes in principal stress magnitudes (left) and directions (right) at Cell 4 located at the main roof coarse-grained sandstone of the pillar. On the horizontal axis, positive numbers indicate the cell is outby the longwall face, and negative numbers indicate the cell is inby the longwall face. The dotted line indicates the start of the longwall face advance.
Fig. 15. Influence of distance on mining-induced side abutment pressure. (a) Δσ 1 , (b) Δσ 2 , and (c) Δσ 3 . The curve numbers indicate the distance from the longwall face to the site location of the stress meters. Positive values indicate the longwall face was outby the site location, and negative values indicate the longwall face was inby the site location.
was 100 m away.18 At a mining depth of 150 m in Emerald mine, front abutment pressure was observed when the longwall face was 33 m away.19 In the present study, the depth of the coal seam was 368–502 m, and there was front abutment pressure when the longwall face was 120 m away. Despite the scatter of data that may be due to the local
Silesian Coal Basin at a mining depth of 740 m, an obvious increase in the front abutment pressure was observed when the longwall face was 200 m away and a larger increase was observed when the longwall face was 40 m away. In the Tashan coal mine, China with a overburden depth of 300–500 m, abutment pressure was observed when the longwall face 9
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International Journal of Rock Mechanics and Mining Sciences 124 (2019) 104141
geological and geotechnical conditions, it is found that the mining depth is a key factor controlling the range of mining-induced stress. The greater the depth, the greater the range of the mining-induced stress. After a coal seam is mined out, the caving of the immediate roof and fracturing of the overburden rocks is a dynamic process that occurs as the longwall face continues to advance, resulting in a gradual change in the mining-induced stress applied to the unmined-out area, i.e., the pillar and entry that will serve the next longwall panel. The monitoring of changes in principal stresses in the present study showed distinct stress changes after the longwall face passed (see Figs. 8–10, 12, and 14), these were a direct indication of the gradual fracturing of the over burden rocks. This finding was similar to a result reported in the liter ature18 that the field monitored pillar stress continued to increase as the longwall face passed by a monitoring station and reached a peak value when the longwall face was 360–379 m away. The continuous advance of the longwall panel seemed to play a role in the gradual fracturing of overburden rocks, as suggested by the distinct periodic change in the principal stresses observed by all five cells. This periodic characteristic of the overburden rock fracturing above the back of the gob plays an important role in the stress condition of the next longwall panel and the surface subsidence resulting from longwall mining.20 Guo et al.21 found that fracturing can reach all the way to the surface when the longwall width-to-depth ratio is greater than 0.75. In the present study, the width of the longwall panel was 180 m, and the depth was 368–502 m, giving a longwall width-to-depth ratio of 0.36–0.49. This suggests that fracturing of the overburden does not reach the ground surface if Guo’s result applies to the present case. The fracturing pattern and process of the overburden varies depending on the location about the panel, leading to a non-uniform distribution of the fractured zone and abutment pressure across the panel width.1 It is clear from Fig. 15 that the abutment pressure is substantially different at different locations from the longwall panel. The greatest increase in the maximum principal stress occurred at Cell 1, which was closest to the panel, suggesting that the peak side abutment pressure occurred in the periphery of the gob with a range of around 11 m. This finding provides fundamental data for the implementation of stress relief measures including blasting and hydraulic fracturing.22,23
maximize the efficiency of a multiple-entry longwall system in terms of ventilation, ground control, economic, and safety considerations. Acknowledgments This work has been supported by the National Key Research and Development Program of China (grant no. 2017YFC0603003). Appendix A. Supplementary data Supplementary data to this article can be found online at https://doi. org/10.1016/j.ijrmms.2019.104141. References 1 Peng S. Coal Mine Ground Control. 2rd ed. New York: Wiley; 1986. second ed. 2 Whittaker BN, Singh RN. Evaluation of the design requirements and performance of gate roadways. Min Eng. 1979;138:535–548. 3 Majdi A. The Stability of Face Access Tunnel in the Deep Soft Rocks of Coal Mining. 1988. 4 Sheorey PR. Design of coal pillar arrays and chain pillars. In: Fairhurst C, ed. Analysis and Design Methods. Oxford: Pergamon; 1993:631–670. https://doi.org/10.1016/ B978-0-08-040615-2.50030-7. 5 Yang W, Lin B, Qu Y, et al. Stress evolution with time and space during mining of a coal seam. Int J Rock Mech Min Sci. 2011;48(7):1145–1152. https://doi.org/ 10.1016/j.ijrmms.2011.07.006. 6 Khanal M, Adhikary D, Balusu R. Evaluation of mine scale longwall top coal caving parameters using continuum analysis. Min Sci Technol China. 2011;21(6):787–796. https://doi.org/10.1016/j.mstc.2011.06.027. 7 Gao F, Stead D, Coggan J. Evaluation of coal longwall caving characteristics using an innovative UDEC Trigon approach. Comput Geotech. 2014;55:448–460. https://doi. org/10.1016/j.compgeo.2013.09.020. 8 Gao F, Stead D, Kang H, Wu Y. Discrete element modelling of deformation and damage of a roadway driven along an unstable goaf — a case study. Int J Coal Geol. 2014;127:100–110. https://doi.org/10.1016/j.coal.2014.02.010. 9 Amadei B, Stephansson O. Rock Stress and its Measurement. London: Chapman & Hall; 1997. 10 Mills KW, Pender MJ. A soft inclusion instrument for in situ stress measurement in coal. In: International Society for Rock Mechanics and Rock Engineering; 1986. http s://www.onepetro.org/conference-paper/ISRM-IS-1986-025. Accessed March 4, 2019. 11 Kang HP, Lin J, Fan MJ. Investigation on support pattern of a coal mine roadway within soft rocks — a case study. Int J Coal Geol. 2015;140:31–40. https://doi.org/ 10.1016/j.coal.2015.01.003. 12 Peng SS. Ground Control Failures: A Pictorial View of Case Studies. West Virginia University; 2007. 13 Kang H, Lou J, Gao F, Yang J, Li J. A physical and numerical investigation of sudden massive roof collapse during longwall coal retreat mining. Int J Coal Geol. 2018;188: 25–36. https://doi.org/10.1016/j.coal.2018.01.013. 14 Brauner G. Rockbursts in Coal Mines and Their Prevention. Rotterdam: A.A. Belkema; 1994. https://www.osti.gov/etdeweb/biblio/143105. Accessed October 9, 2019. 15 Konicek P, Waclawik P. Stress changes and seismicity monitoring of hard coal longwall mining in high rockburst risk areas. Tunn Undergr Space Technol. 2018;81: 237–251. https://doi.org/10.1016/j.tust.2018.07.019. 16 Pt� a�cek J. Rockburst in ostrava-karvina coalfield. Procedia Eng. 2017;191:1144–1151. https://doi.org/10.1016/j.proeng.2017.05.289. 17 Chen J, Mishra M, Zahl E, Dunford J, Thomas R. Longwall mining-induced abutment loads and their impacts on pillar design and entry stability. In: Proceedings of the 21st International Conference on Ground Control in Mining. WV, USA: Morgantown; 2002: 11–17. 18 Yu B, Zhang Z, Kuang T, Liu J. Stress changes and deformation monitoring of longwall coal pillars located in weak ground. Rock Mech Rock Eng. 2016;49(8): 3293–3305. https://doi.org/10.1007/s00603-016-0970-8. 19 Barczak TM, Tadolini SC, Zhang P. Evaluation of support and ground response as longwall face advances into and widens pre-driven recovery room. In: Proceedings of the 26th International Conference on Ground Control in Mining. WV: Morgantown; 2007:160–172. 20 Jeran PW, Trevits MA. Timing and Duration of Subsidence Due to Longwall Mining. U.S. Department of the Interior; 1995. https://www.cdc.gov/niosh/mining/works/cov ersheet1746.html. Accessed October 11, 2019. 21 Guo H, Adhikary DP, Gabeva D. Hydrogelogical Response to Longwall Mining. 2007: 148. 22 Gale WJ, Nemcik J, Upfold RW. Application of stress control methods to underground coal mine design in high lateral stress fields. In: Proc 6th International Conference on Rock Mechanics. vol. 25. 1987:897–900. https://doi.org/10.1016/ 0148-9062(88)91803-7. Montreal. 23 Kang H, Lv H, Gao F, Meng X, Feng Y. Understanding mechanisms of destressing mining-induced stresses using hydraulic fracturing. Int J Coal Geol. 2018;196:19–28. https://doi.org/10.1016/j.coal.2018.06.023.
5. Conclusions The following conclusions are drawn from the field monitoring data, which provide valuable information to better understand the load transfer mechanics associated with longwall retreat mining. Mining depth is a key factor controlling the range of mining-induced stress. The greater the depth, the greater the range of the mininginduced stress. The influence zone of the front abutment load on the trail entry reached as far as 120 m outby the longwall face during panel mining. The influence of the front abutment pressure, however, was not significant until the longwall face was 20 m inby. As the longwall face continued to approach and pass, there was a dramatic increase in the deformation of the trail entry. Owing to the high abutment pressure, bed separation could occur in the competent fine-grained sandstone that was 7.4 m deep into the trail entry roof. The first support pattern consisting of high-strength rock bolts and cable bolts installed with a W-shaped steel strap and welded screen was found to be effective in sustaining the integrity of the trail entry. Periodic collapse of the main roof above the mined-out area continued when the longwall face had advanced 120 m away. The peak side abutment pressure occurred in the periphery of the gob with a range of around 11 m. The results of the present study improve our understanding of the local transfer mechanics associated with longwall coal mining. Such understanding allows the development of design criteria that will help
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