Numerical Simulation on Supporting of Large Section Open-off Cut and Its Stability

Numerical Simulation on Supporting of Large Section Open-off Cut and Its Stability Weijian Yu Hunan Provincial Key Laboratory of Safe Mining Techniques of Coal Mines,Hunan University of Science and Technology,iangtan,Hunan 411201,China e-mail:ywjlah@163.com Shaohua Du School of Energy and Safety Engineering,Hunan University of Science and Technology,iangtan,Hunan 411201,China e-mail: 2465492388@qq.com ABSTRACT To solve the deformation problem of large section open-off cut of loose coal seam, numerical simulation, field test and deformation monitoring were used to study the roadway supporting of the 212 face cut in one coal mine. First, according to the actual geological conditions, the method of the engineering analogy and theoretical calculation were preliminarily determined the forms of combined supporting and implementation about the roadway supporting parameters, the anchor cable, the w type steel and metal mesh, etc. Then, it could be to determine the original state and the calculation parameters of excavation and supporting by using FLAC 2D software after considering the size effect of excavation, these four steps were the left half of cut 3.4m wide roadway, the left side of the roadway bolt and cable anchor supporting, the bolt and cable anchor supporting of the right half of the cut roadway and the right half of the cut roadway, which were calculated and analyzed separately. Finally, the supporting was used in the field, and the monitoring results shown that the maximum relative displacement of the roof and floor was 123mm, and the maximum value of two sides move quantity was 26mm. The deformation tended to be stable, and the numerical results are consistent with field observations. KEYWORDS: seam roadway; fully mechanized caving face; open-off cut; support design; numerical simulation - 1619 -

Vol. 19 [2014], Bund. H 1620 INTRODUCTION For large section in deep seam roadway support, it generally used bolt and anchor combined supporting forms (1,2). However, due to the loose seams, fracture surface and span a larger, supporting arguments unreasonable, and other engineering complex geological conditions and other reasons, cut roadway often occurred instability phenomenon before coal mining. Therefore, some researchers and engineers specifically studied and optimized on supporting of large section open-off cut. He (3,4) used of numerical simulation software on the original support programs and currently support programs to conduct the related numerical simulation analysis and comparison, and his result reveals that the use of pressures composite truss anchor support system to controls large section open-off cut country rock was safe and reliable; analyzing the roof lithology of open-off cut, Wei (5) used bolt (cable) supporting theory and mechanism to design supporting parameters of large section open-off cut of secondary expansion, and determined the open-off cut in secondary roof to expand the use of the 20 mm 4 000 mm thread steel anchor and 21.6 mm 12 000 mm strand anchor reinforcement; Yin (6) revealed the main cause of the failure supporting of large cross section open-off cut with watery broken roof, using stepwise combined support, Roof hydrophobic, water sentinel grouting and sealing technology in harmony, stability Cut off status of the field observations; Zhang (7) combined with FLAC 3D software to analyze its rock yield failure after calculation and analysis, top floor, roadway's sides closer to the amount, etc., then reasonable support technology solutions of no.2 seam face cut come out; with using engineering analogy method that could determine supporting program, Nan (8) used FLAC 3D software in different Supporting forms to conduct simulation, and study the main factors, and so on. The above research results in the actual production process played a very important role in guiding. However, the length of the new tunneling roadway in Chinese state-owned coal mines each year reached up to 8000km, in which more than 80% was coal roadway, and most coal roadway was in a complex application environment, the mechanical mechanism is different, many of the factors affecting instability. Therefore, the stability of loose coal seam large section roadway under complex conditions had been the focus of mining technology and engineering workers. On the basis of a 212 caving fully mechanized face specific engineering geology and supporting forms, this paper used numerical simulation stability analysis, and in the support program live rock deformation observation, etc., provided a scientific basis for the actual production. THE GEOLOGICAL CONDITIONS AND SUPPORTING OF CAVING FULLY MECHANIZED FACE CUT For a coal mine 212 face cut, its length was 164m, seam thickness was 7.11m, it contained two layers of stone, and the upper part was about eight inches stone, whose thickness was about 0.09m form roof; the lower part was 1.9m refine stone, whose thickness was 0.47m. The middle

Vol. 19 [2014], Bund. H 1621 part of the coal seam was loose, with the excavation dig with risk. The immediate roof lithology was black silty mudstone, and its thickness was 1.1m. The main roof lithology was limestone; its thickness was 10.33m. Roadway was about 430 ~ 580m from surface. The roadway was rectangular in shape, along the floor of coal seam tunneling, width 6.5m, height 2.7m, width 6.8m height 2.8m tunneling, and excavation. Through the engineering analogy and theoretical calculation, preliminary determined the roadway support parameters (Figure 1). Bolt, cable, w steel and metal mesh combined supporting were used in roof using. Roof bolt: The tunnel roof with diameter of 22mm, length of 2.6m high strength steel bolt, the double speed of resin anchoring agent. Anchorage length was 1.0m, force value was not less than 70kN, and the distance of bolt nut tightening torque was not less than 120N m. Inter-row spacing are 0.8m, rectangular layout; W steel band: W steel band with B220 3800 3mm, B220 3400 3mm two block, spacing of 0.8m; (3) metal mesh: 10 # lead wire metal mesh used grid of 60 60mm diamond wire mesh, and all covered; Pre-stress anchor: With small diameter prestress anchor, 1860 level of 7 strands of steel strand. Anchor length 7.3m, QLM15-1-type anchors, with 14 # channel steel joists, long 2.7m, and parallel to the roadway layout. There were two cables to fix each root beam, and their span was 1.6m. Along the cut section in the roof were evenly arranged four rows of anchor cable, row spacing was 1.4m, and its shape was rectangular. Anchor bearing capacity was 230kN, preload 120 ~ 150kN. (a) Plans (b) cross-section Figure 1: Support parameters of a mine's 212 large section cut Metal mesh bolting form were used in two sides bolting. It need the ordinary steel bolt that diameter was 20mm and length was 2.2m, ordinary steel, medium resin anchoring agent, and 600 150 80mm wood pallets that anchoring length was 0.8m. Anchoring force value was not less than 50kN, tightening torque was not less than 60N m, inter-row spacing was 0.7 0.8m, and its shape was rectangular. The metal mesh supporting cons stented with the roof.

Vol. 19 [2014], Bund. H 1622 NUMERICAL ANALYSIS ON ANCHOR SUPPORTING OF LARGE LECTION OPEN-OFF CUT Numerical calculation model In order to consider the boundary effect on the calculation results, the two left and right boundary of calculation model used the stress boundary, and the lower boundary used displacement boundary. Self-weight stress on the boundary was caused by the overlying strata. According to the description of mining engineering geological data, we could determine Position and lithology of coal and rock strata roadway. On the basis of the coal seam and roof and floor rock physical mechanics test and in-situ stress test results, we could also determine simulation calculation parameters of original rock stress field of rock and the simulation of excavation and support parameters of engineering rock mass. In order to calculate the original rock stress field, indoor rock mechanics test results should be reduced appropriately to simulate calculation parameters of the original rock stress field of mining. Obviously, laboratory parameters were not the actual parameters of rock mass, and the main reason was the size effect of rock mass; at the same time, it was also because of sample tests on rock disturbance and the influence of other factors. Simulation calculation parameters of the initial stress field were shown in Table 1. location and lithology location Table 1: calculation parameters of 212 caving fully mechanized face cut in the original rock lithology Volume Density γ/kg/m 3 Tensile strength Shear strength parameters Angle of Cohesion internal friction Modulu s Deformation parameters Poisson's ratio bulk modulus shear modulus R /MPa C/MPa t φ ( o ) E/GPa μ B/GPa G/GPa Main roof Dark gray limestone 2680 14.2 19.0 53.0 9.78 0.263 6.88 3.87 Immediate Black roof mudstone 2080 5.0 5.88 47.0 2.00 0.195 1.09 0.84 Soal seam Coal seam 1460 3.2 3.31 42.0 1.48 0.275 1.09 0.58 floor Sandy mudstone 2640 7.2 9.19 51.1 5.68 0.278 4.26 2.22 Because of coal mining, roadway and stope on a range of original rock produced different degrees disturbance. Meanwhile, the jointed rock mass and the role of groundwater caused that the physical and mechanical properties of the rock mass produce different degrees reduction. It was necessary that the original rock parameters were reduced as the parameters of coal and rock mass. According to the engineering experience, the reduced factor was given in Table 2. Engineering rock parameters that were determined by reduction factors based on the table were listed in Table 3.

Vol. 19 [2014], Bund. H 1623 Table 2: calculation parameters and reduced factors of 212 caving fully mechanized face roadway location and lithology Shear strength parameters Volume Tensile location lithology Density strength Cohesion Main roof Dark gray limestone Angle of internal friction Deformation parameters Modulus Poisson's ratio 0.98 0.10 0.15 0.85 0.95 1 Immediate roof Black mudstone 0.95 0.00 0.05 0.70 0.90 1 Soal seam Coal seam 0.90 0.00 0.00 0.60 0.85 1 floor Sandy mudstone 0.95 0.05 0.05 0.70 0.90 1 location and lithology location Table 3: calculation parameters of 212 caving fully mechanized face roadway excavation and supporting lithology Volume Density γ/kg/m 3 Tensile strength Shear strength parameters Angle of Cohesion internal friction Modulus Deformation parameters bulk Poisson' shear modulu s ratio modulus s R /MPa C/MPa t φ ( o ) E/GPa μ B/GPa G/GPa Main roof Dark gray limestone 2626 1.42 2.85 45 9.29 0.263 6.53 3.68 Immediate Black roof mudstone 1976 0.00 0.29 33 1.80 0.195 0.98 0.75 Soal seam Coal seam 1314 0.00 0.00 25 1.26 0.275 0.93 0.49 floor Sandy mudstone 2508 0.36 0.46 36 5.11 0.278 3.84 2.00 The calculation model and the anchor net support parameters In order to simulate the large section cut excavation and support process, the calculation parameters in Table 1 was used to simulate rock original rock stress field. On this basis, the excavation process and support parameters of roadway were taken into account to make simulation and analysis. Table 2 was calculation model of large section cut supporting numerical simulation. According to the anchor net support parameters and bolt and anchor cable supporting form shown in the figure 1, and considering the support spacing, processing of the plane strain problem of equivalent, we could get the parameters of support parameters of bolt. Bolt supporting parameters and properties and equivalent calculation parameters were shown in Table 4 and Table 5.

Vol. 19 [2014], Bund. H 1624 4.000 21-Jan-07 1:48 step 4915-2.833E+00 <x< 5.383E+01-9.833E+00 <y< 4.683E+01 3.000 0 1E 1 Grid plot 0 1E 1 Fixed Gridpoints -direction Y Y-direction 1.000 B Both directions Net Applied Forces max vector = 1.020E+07 0 2E 7 BYYYYYYYYYYYYYYYYYYYYYYYYYYYYYYYYYYYY YYYYYYYYYYYYYYYYYYYYYB 0.000 0.500 1.500 2.500 3.500 4.500 Location Figure 2: numerical model of 212 caving fully mechanized face cut anchor net supporting Table 4: properties and parameters of 212 caving fully mechanized face cut anchor net supporting Length l / m Diameter d/mm Spacing a / m Row spacing t / m Anchorage length l m /m Modulus of elasticity E/GPa Anchoring yield value yi / MN Mortar compressive strength /GPa The shear modulus of mortar /GPa Roof 2.6 22 0.8 0.8 0.8 98.6 0.548 20.0 9.0 Ribs 2.2 20 0.7 0.80 0.6 98.6 0.548 20.0 9.0 Anchor 7.3 15.24 1.6 1.4 2.0 98.6 0.548 20.0 9.0 location Table 5: calculation parameters of 212 caving fully mechanized face cut anchor net supporting Shear stiffness Kb/ GPa The actual calculation parameters Shear strength Sb/ kn m -1 Modulus of elasticity E/GPa Yield value Yie/ MN Cross-sectional area Ar/ m 2 Shear stiffness Kb/ GPa Parameter calculation of equivalent plain strain Shear strength Sb/ kn m -1 Modulus of elasticity E/GPa Yield value Yie/ MN Crosssectional area Ar/ m 2 roof 23.4 691 98.6 0.548 3.8 10-4 29.25 864 123 0.685 3.8 10-4 ribs 16.8 628 98.6 0.548 3.14 10-4 21.0 785 123 0.685 3.14 10-4 anchor 9.2 729 98.6 0.548 1.82 10-4 6.6 521 70.4 0.391 1.82 10-4

Vol. 19 [2014], Bund. H 1625 The roadway excavation and supporting numerical simulation and analysis The first calculation step: Firstly, it is excavation cut left half 3.4m wide roadways; Second calculation step: Bolt and anchor supporting was applied to the left side of the roadway; Third calculation step: Excavation cut left half roadways; Fourth calculation step: Bolt and anchor support was applied to the right half of the roadway. (1) The first calculation step (the left side of the roadway excavation) After the left side cut excavation, figure 3 was the roadway deformation and displacement vector map. This figure shown that if the anchor net support had been not timely after excavation, the roadway would have a large deformation. At this time, roof displacement had reached 5.8cm, floor displacement was close to 4cm, therefore, relative displacement of roof and floor was close to 10cm, and the displacement was enough to make the roadway surface occur drum crack damage. So, after the first excavation, the implementation of the cut secondary excavation was firstly applied steel mesh to the excavation face so as to form a protective layer; at the same time, also bolt and cable anchor supporting should be used to minimize harmful loose caused by excessive deformation of surrounding rock as early as possible. The distribution of plastic zone after the left side cut excavation was shown in Figure 4. According to this figure, there was a large range of plastic yielding zone in surrounding rock because the roadway was not supported. 21-Jan-07 2:07 step 10039 1.700E+01 <x< 3.400E+01 1.100E+01 <y< 2.550E+01 0 5E 0 Grid plot 0 5E 0 Displacement vectors max vector = 5.803E-02 0 1E -1 2.400 2.200 1.800 1.600 1.400 1.200 1.800 2.200 2.400 2.600 2.800 3.000 3.200 Figure 3: The roadway deformation and displacement vector map after the left side cut excavation

Vol. 19 [2014], Bund. H 1626 4.000 21-Jan-07 2:30 step 10381-2.833E+00 <x< 5.383E+01-9.833E+00 <y< 4.683E+01 3.000 0 1E 1 Cable plot Plasticity Indicator * at yield in shear or vol. elastic, at yield in past o at yield in tension 1.000 0.000 0.500 1.500 2.500 3.500 4.500 Figure 4: the cut excavation left anchor Shotcrete roadway displacement map (2)Second calculation step (anchor net supporting) Displacement vector map after anchor net supporting were shown in Figure 5. With timely anchor net supporting, roadway deformation was relatively small, and the maximum displacement was less than 5cm. From the figure, because of the right side of left roadway without supporting, it could be found that the deformation was larger than the left side and the bolt and cable anchor supporting plays a very significant role on the stability of surrounding rock. Obviously, because of the right side of surrounding rock without anchor net supporting, there would also be significantly increased displacement in the right side of the roadway and the right side of roof and floor rock. Therefore, it was necessary to strengthen the timeliness and integrity of roadway supporting so as to maintain the stability of roadway. The maximum principal stress map of surrounding rock after bolt mesh supporting is shown in Figure 6. As shown in this figure, there was a low stress area in the roadway floor, and indicated that surrounding rock was in the tensile stress state or closed to the tensile stress state of near the bottom. Unlike the floor, exerted the bolt and anchor supporting, the roof and the left side of the surrounding rock increased the rock's own strength, therefore, and also improved the bearing capacity of surrounding rock. There was high pressure stress that was more conducive to stability of the stress state. A plastic zone distribution map of the surrounding rock after the left half of anchor net supporting is shown in Figure 7. Thus, because of the right side of roadway without supporting, the plastic yield zone existed in a wide range of the right side of surrounding rock; Compared with the left of surrounding rock and because of bolt and cable supporting role, most of the region had recovered the elastic state even though there were a plastic yield area, and improved the self bearing capacity of surrounding rock.

Vol. 19 [2014], Bund. H 1627 2.200 21-Jan-07 2:32 step 10014 1.800E+01 <x< 3.200E+01 1.100E+01 <y< 2.300E+01 0 2E 0 Grid plot 0 2E 0 Displacement vectors max vector = 4.500E-02 1.800 1.600 0 1E -1 1.400 1.200 1.900 2.100 2.300 2.500 2.700 2.900 3.100 Figure 5: the displacement vector map after anchor net supporting of the left half of roadway cut Figure 6: the maximum principal stress map of surrounding rock after anchor net supporting of the left half of roadway cut

Vol. 19 [2014], Bund. H 1628 4.000 21-Jan-07 2:32 step 10014-2.833E+00 <x< 5.383E+01-9.833E+00 <y< 4.683E+01 3.000 0 1E 1 Cable plot Plasticity Indicator * at yield in shear or vol. elastic, at yield in past 1.000 0.000 0.500 1.500 2.500 3.500 4.500 Figure 7: the plastic zone distribution map of surrounding rock after the left half of anchor net supporting of the left half of roadway cut (3)Third calculation step (excavation on the right side of the roadway, and shot Crete supporting) Displacement vector map after the right side of the roadway excavation is shown in Figure 8. Thus, the maximum deformation of surrounding rock was close to 7cm. Maximum principal stress isocline map of third calculation step of surrounding rock is shown in Figure 9. Therefore, there was a larger range low stress area in the roadway floor. In contrast, because bolt and cable anchor supporting was applied in the left side and roof, greatly improves the bearing capacity of surrounding rock and the rock was more conducive to the stability of compressive stress state. Plastic zone map of the calculation step of surrounding rock is shown in Figure 10. Thus, the right not to anchor support, there was a wide range of plastic zone on the right side of roadway surrounding rock. In comparison, the scope of plastic zone in the left side of surrounding rock was relatively small, and because of the anchor net supporting, part of the plastic zone of surrounding rock intensity increased so as to restore to the elastic state. Obviously, the original tensile stress damage zone existed in the roof and the anchor net supporting, made it no longer exist so as to improve the stability of surrounding rock effectively.

Vol. 19 [2014], Bund. H 1629 2.200 21-Jan-07 2:33 step 10321 1.800E+01 <x< 3.200E+01 1.100E+01 <y< 2.300E+01 0 2E 0 1.800 Grid plot 0 2E 0 Displacement vectors max vector = 6.833E-02 1.600 0 2E -1 1.400 1.200 1.900 2.100 2.300 2.500 2.700 2.900 3.100 Figure 8: The displacement vector map after the right side of the roadway cut excavation Figure 9: The maximum principal stress isoline map after the right side of the roadway cut excavation

Vol. 19 [2014], Bund. H 1630 4.000 21-Jan-07 2:33 step 10321-2.833E+00 <x< 5.383E+01-9.833E+00 <y< 4.683E+01 3.000 0 1E 1 Cable plot Plasticity Indicator * at yield in shear or vol. elastic, at yield in past 1.000 0.000 0.500 1.500 2.500 3.500 4.500 Figure 10: The plastic zone map of surrounding rock after the right side of the roadway cut excavation (4)Fourth calculation step (the final results of anchor net supporting) Displacement vector map of the last part of roadway is shown in Figure 11. Therefore, after step excavation and supporting, the deformation of the surrounding rock could be effectively controlled, and the maximum displacement vector was less than 7cm, which was close to the value measured displacement. For large section coal roadway, it shown that the implementation of long cable control technology and supplementing with comprehensive support technical of short bolt and metal net were able to meet the requirements of the stability of roadway. Plastic zone of roadway final state is shown in Figure 12. Therefore, because of the long cable anchor supporting, the plastic zone of the surrounding rock was significantly reduced. The plastic zone of roof surrounding rock completely recovered, and the plastic zone of the right side did not exist. Although the plastic zone of the left side of surrounding rock also was reduced, there was still a shear failure zone in the upper left and lower. Seen from this, after applying long anchor supporting to the roof, both sides still exist a shear failure zone that was the weak area of roadway stability. Therefore, appropriately increased the length of the roof wings and both sides of the upper bolt, it was necessary to prevent the shear failure of roadway in the corners.

Vol. 19 [2014], Bund. H 1631 2.200 21-Jan-07 2:34 step 13786 1.800E+01 <x< 3.200E+01 1.100E+01 <y< 2.300E+01 0 2E 0 Grid plot 0 2E 0 Displacement vectors max vector = 6.828E-02 1.800 1.600 0 2E -1 1.400 1.200 1.900 2.100 2.300 2.500 2.700 2.900 3.100 Figure 11: the displacement vector map after the excavation supporting of cut 4.000 21-Jan-07 2:34 step 13786-2.833E+00 <x< 5.383E+01-9.833E+00 <y< 4.683E+01 3.000 0 1E 1 Cable plot Plasticity Indicator * at yield in shear or vol. elastic, at yield in past 1.000 0.000 0.500 1.500 2.500 3.500 4.500 Figure 12: The plastic zone map of surrounding rock after the excavation supporting of roadway DEFORMATION OBSERVATION AND ANALYSIS OF THE SCENE According to the geological conditions of the mine, a typical roadway was choose to observe the deformation. With building a roadway observatory each forward 50m, two observation section were selected up in each observatory. A total of 5 observation stations were used to monitor the relative displacement of roof and floor. The deformation monitoring data are shown in Table 6. It is shown respectively the displacement curve of both sides and the top and bottom of the typical observation roadway Figure 13 and Figures 14. Thus, the relative displacement of the top and bottom of roadway were respectively 56mm, 45mm, 16mm, 93mm, 123mm; ten days ago, the displacement of tunneling effect stage were respectively 40mm,

Vol. 19 [2014], Bund. H 1632 27mm, 10mm, 60mm, 91mm; the maximum displacement velocity were respectively 10mm/d, 7mm/d, 4mm/d, 10mm/d, 56mm/d. The change of the last two observation stations was bigger. Because there were a larger cracks and a larger soft coal chute within the scope of these two observatories, seam joints were well developed, and coal body was loose. However, from the perspective of the overall monitoring results, it is can conclude that the deformation of roadway tended to be stable. In addition, according to the observation data, a larger range of time and scope of the relative displacement of the roof and floor appeared in the roadway excavation just 10 days, and its distance was 40m or so in front of the dig working face. The maximum displacement was 123mm, and the maximum displacement speed was 56mm/d. Two-sided displacement amount: the relative displacement of two sides of roadway was not obvious, and the maximum displacement is 26mm. The displacement speed was 0.87mm/d, and the maximum displacement speed is 8mm/d. The numerical simulation results show that the maximum displacement of roadway was less than 70mm and basically cons stented with the field observation data without considering the influence of joint fissure. Seen from the results of field application, the supporting could ensure that the working face was stable before the formal mining, and also ensure the practical requirements of the equipment, the pedestrian and the ventilation. Figure 12: The displacement curve of the roof and floor Figure 13: The displacement curve of both sides

Vol. 19 [2014], Bund. H 1633 Table 6: The surface displacement of roadway Observation points The first observation point The third observation point The fifth observation point 0~10 11 to 30 days 30 days later total 0~10 11 to 30 days 30 days later total 0~10 11 to 30 days 30 days later total Project Surface displacement of roadway The amount of displacement (mm) The average speed (mm/d) Maximum speed (mm/d) Observation points Project Surface displacement of roadway The amount of displacement (mm) The average speed (mm/d) Maximum speed (mm/d) roof 40 4 10 roof 27 2.7 7 0~10 ribs 10 1 1.5 ribs 8 0.8 2 roof 12 1.2 3 11 to roof 13 0.65 3 ribs 0 0 0 The second 30 observation days ribs 2 0.01 1 roof 4 0.2 1 point 30 roof 5 0.07 3 ribs 0 0 0 days later ribs 0 0 0 roof 56 1.8 10 roof 45 1.14 7 total ribs 10 0.03 0.05 ribs 10 0.27 2 roof 10 1 4 roof 60 6 10 0~10 ribs 3 0.3 2 ribs 16 1.6 8 roof 4 0.2 1 11 to roof 22 2.2 7 ribs 1 0.05 1 The fourth 30 observation days ribs 6 0.6 2 roof 2 0.1 2 point 30 roof 11 1.1 4 ribs 0 0 0 days later ribs 2 0.2 1 roof 16 0.4 4 roof 93 3.1 10 total ribs 4 0.01 2 ribs 24 0.8 8 roof 91 9.1 56 ribs 21 2.1 5 roof 24 2.4 5 ribs 4 0.4 2 roof 8 0.8 3 ribs 1 0.1 1 roof 123 4.1 56 ribs 26 0.87 5 CONCLUSION By using the numerical analysis method for the simulation of roadway excavation and supporting of large section cut, it revealed the variation rule of displacement, stress and plastic zone of the large section roadway in the different stage of construction. From the analysis result, it is can see that the Stepwise excavation of roadway and the combined supporting of bolt, anchor cable, W type steel and metal mesh, etc. Deformation could controlled the deformation of roadway surrounding rock of large section cut effectively, and the maximum displacement was less than 70mm and could meet the requirements of the stability of roadway.

Vol. 19 [2014], Bund. H 1634 ACKNOWLEDGEMENTS This work was financially supported by the National Natural Science Foundation of China (51104063, 51374105, 51374106, and 51174086) and Scientific Research Fund of Hunan Provincial Education Department (12cy013). REFERENCES 1. Weijian YU, Weijun WANG, Guohua WEN, (2012). Deformation mechanism and control technology of coal roadway under deep well and compound roof [J]. Chinese Journal of Geotechnical Engineering,34(8):1501-1508. 2. Weijian YU, Weijun WANG, Nong ZHANG, (2012). Study of global deformation and control of a thick, layered compound roof in a deep well [J]. Journal of China University of Mining & Technology,41(5):725-732. 3. Fulian He, Jianping Li, Hongjun Jiang, (2012). The application of pre-stress compound truss cable in set-up room with large cross section thick coal roof [J],China Mining Magazine,21(8):82-85 4. Fulian He, Dongping Yin, Hong Yan, (2010). Study on the coupling system of high pre-stress cable truss and surrounding rock on a coal roadway [A],Rock Stress and Earthquakes[C], Netherlands: CRC PRESS 5. Shiyi WEI. (2013). Study on Support Parameters Optimization of Large Cross Section Open- ff Cut with Broken Roof in Deep Underground Mine [J],Coal Science and Technology,41(4):28-32 6. Dajun YIN. (2013). Study on Surrounding Rock Control Technology of Large Cross Section Open-off Cut with Watery Broken Roof [J],Coal Science and Technology,41(5):35-38 7. Airong Zhang, Kai Wang. (2013). Numerical Simulation of Mining Face Cut Supporting Technology [J],Journal of Taiyuan University of Technology,41(5):35-38 8. Qinan YANG. (2012). Numerical simulation of compound roof control at coal roadway with large cross-section based on FLAC3D [J],Journal of Liaoning Technical University(Natural Science),31(4):461-465 2014, EJGE