Failure mechanism and stability control technology of rock surrounding a roadway in complex stress conditions Yu Yang a,b,⇑ , Bai Jianbiao a,b , Chen Ke c , Wang Xiangyu a,b , Xiao Tongqiang a,b , Chen Yong a,b a School of Mines, China University of Mining & Technology, Xuzhou 221116, China b State Key Laboratory of Coal Resources and Mine Safety, China University of Mining & Technology, Xuzhou 221008, China c Baoji Coal Industry Corporation Ltd., Baoji 721000, China article info Article history: Received 14 August 2010 Received in revised form 9 September 2010 Accepted 10 October 2010 Available online 12 June 2011 Keywords: Both sides mining Numerical simulation FLAC 3D Grouting reinforcement Support parameters abstract To solve the problem of supporting three downhill coal structures in the Yongan Coal Mine of Shanxi Jincheng, we studied the regular development of stress and plastic zones and characteristics of deforma- tion of rock surrounding roadway groups after a period of roadway driving, mining one side as well as mining both sides, we used FLAC 3D for our numerical and theoretical analyses. Field test were carried out, where we revealed the deformation mechanism of roadways and its coal pillars in complex stress conditions. We proposed a roadway stability control technology using backwall grouting with high-water rapid hardening material and combined support with bolt and cable anchoring after mining both sides. Our field practices showed that deformation of rock surrounding roadways can be controlled with this technology. Ó 2011 Published by Elsevier B.V. on behalf of China University of Mining & Technology. 1. Introduction In China, 70–80% of roadways in mines are affected by working faces in the process of mining coal [1–3]. In recent years, with sub- stantial increases in mining scale, intensity, output and mining depth, the maintenance of roadways, under dynamic pressure, is becoming increasingly more difficult. Of late, this kind of roadway has become the bottleneck by restricting more intensive coal mine output. Roadways under dynamic pressure are characterized by damage to their original state of equilibrium, expanding fractured surfaces, large deformations and surrounded by loose broken rock. These conditions make their maintenance extremely difficult [2–5]. Therefore, research into the failure mechanism of the surrounding rock and of coal pillars, given the effect of disturbance stress, is ur- gently needed for safe working conditions in mines, as is developing a control technology to maintain stability for this kind of roadway. As well, such measures should improve economic benefits. 2. Situation of test roadway The #3 coal seam, being mined in the Yongan Coal Mine of Shanxi Jincheng, is of a simple structure and high hardness. The seam is 700 m deep, 6.26 m thick with a dip angle of 4°. The roof of this coal seam, with a lithology of black gray siltstone and mud stone, is 2.54 m thick and caves easily. The floor of the #3 coal seam with a lithology of black gray mud stone is 2.18 m thick and easily musters when forced. The mining height of the fully mecha- nized #3210 working face, part of the lower layers of the #3 coal seam, is 3.2 m. The mining of the upper layers of working face #3209 has been completed. The downhill material, belt and return air lie towards the north of working face #3210 with the #3207 and #3208 goafs to its east. Towards the west, the coal seam has not yet been mined. The boundary of the mine field lies towards the south. The goaf of the Wulimiao Coal Mine lies to the north of these three downhills. There is a 16 m coal pillar between the downhill material and belt and an 18 m coal pillar between the downhill belt and the air return. The spatial relation between the three downhill locations and the working face is shown in Fig. 1. The effects of multiple mines on the three downhill structures are broken surrounding rocks, a caved primary dressed stone arch supporting wall, serious floor heaving and large sectional shrinkage of between 55% and 85%. Maintenance of the three structures is still very poor, although two renovations have been carried out and we used even a double-layered dressed stone arch to support the wall. Damage to the three downhill structures is shown in Fig. 2. 3. Numerical model In order to study the regular development of the stress, the plastic zone and the deformation characteristics of the rock 1674-5264/$ - see front matter Ó 2011 Published by Elsevier B.V. on behalf of China University of Mining & Technology. doi:10.1016/j.mstc.2011.05.024 ⇑ Corresponding author. Tel.: +86 13852004679. E-mail address: [email protected](Y. Yu). Mining Science and Technology (China) 21 (2011) 301–306 Contents lists available at ScienceDirect Mining Science and Technology (China) journal homepage: www.elsevier.com/locate/mstc
Transcript
Mining Science and Technology (China) 21 (2011) 301–306
Contents lists available at ScienceDirect
Mining Science and Technology (China)
journal homepage: www.elsevier .com/locate /mstc
Failure mechanism and stability control technology of rock surroundinga roadway in complex stress conditions
Yu Yang a,b,⇑, Bai Jianbiao a,b, Chen Ke c, Wang Xiangyu a,b, Xiao Tongqiang a,b, Chen Yong a,b
a School of Mines, China University of Mining & Technology, Xuzhou 221116, Chinab State Key Laboratory of Coal Resources and Mine Safety, China University of Mining & Technology, Xuzhou 221008, Chinac Baoji Coal Industry Corporation Ltd., Baoji 721000, China
a r t i c l e i n f o a b s t r a c t
Article history:Received 14 August 2010Received in revised form 9 September 2010Accepted 10 October 2010Available online 12 June 2011
To solve the problem of supporting three downhill coal structures in the Yongan Coal Mine of ShanxiJincheng, we studied the regular development of stress and plastic zones and characteristics of deforma-tion of rock surrounding roadway groups after a period of roadway driving, mining one side as well asmining both sides, we used FLAC3D for our numerical and theoretical analyses. Field test were carriedout, where we revealed the deformation mechanism of roadways and its coal pillars in complex stressconditions. We proposed a roadway stability control technology using backwall grouting with high-waterrapid hardening material and combined support with bolt and cable anchoring after mining both sides.Our field practices showed that deformation of rock surrounding roadways can be controlled with thistechnology.
� 2011 Published by Elsevier B.V. on behalf of China University of Mining & Technology.
1. Introduction
In China, 70–80% of roadways in mines are affected by workingfaces in the process of mining coal [1–3]. In recent years, with sub-stantial increases in mining scale, intensity, output and miningdepth, the maintenance of roadways, under dynamic pressure, isbecoming increasingly more difficult. Of late, this kind of roadwayhas become the bottleneck by restricting more intensive coal mineoutput.
Roadways under dynamic pressure are characterized by damageto their original state of equilibrium, expanding fractured surfaces,large deformations and surrounded by loose broken rock. Theseconditions make their maintenance extremely difficult [2–5].Therefore, research into the failure mechanism of the surroundingrock and of coal pillars, given the effect of disturbance stress, is ur-gently needed for safe working conditions in mines, as is developinga control technology to maintain stability for this kind of roadway.As well, such measures should improve economic benefits.
2. Situation of test roadway
The #3 coal seam, being mined in the Yongan Coal Mine ofShanxi Jincheng, is of a simple structure and high hardness. Theseam is 700 m deep, 6.26 m thick with a dip angle of 4�. The roof
Elsevier B.V. on behalf of China Un
of this coal seam, with a lithology of black gray siltstone andmud stone, is 2.54 m thick and caves easily. The floor of the #3 coalseam with a lithology of black gray mud stone is 2.18 m thick andeasily musters when forced. The mining height of the fully mecha-nized #3210 working face, part of the lower layers of the #3 coalseam, is 3.2 m. The mining of the upper layers of working face#3209 has been completed. The downhill material, belt and returnair lie towards the north of working face #3210 with the #3207and #3208 goafs to its east. Towards the west, the coal seam hasnot yet been mined. The boundary of the mine field lies towardsthe south. The goaf of the Wulimiao Coal Mine lies to the northof these three downhills. There is a 16 m coal pillar between thedownhill material and belt and an 18 m coal pillar between thedownhill belt and the air return. The spatial relation between thethree downhill locations and the working face is shown in Fig. 1.
The effects of multiple mines on the three downhill structuresare broken surrounding rocks, a caved primary dressed stone archsupporting wall, serious floor heaving and large sectional shrinkageof between 55% and 85%. Maintenance of the three structures is stillvery poor, although two renovations have been carried out and weused even a double-layered dressed stone arch to support the wall.Damage to the three downhill structures is shown in Fig. 2.
3. Numerical model
In order to study the regular development of the stress, theplastic zone and the deformation characteristics of the rock
Fig. 1. Spatial relations between three downhill locations and working face. (1)Downhill material; (2) downhill belt; (3) downhill return air; (4) direction of upperlayered working face #3210; (5) upper layered working face #3209; (6) upperlayered working face #3210; and (7) goaf of the Wulimiao Coal Mine.
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surrounding the roadway after the effect of roadway driving andcoal mining, we need to analyze the deformation and failure mech-anism. For that, we have used a large-scale software for geotechni-cal analysis, FLAC3D, in order to establish a three-dimensionalnumerical model. It is a physical elastoplastic model, with a yieldcriterion in Mohr-Coulomb. The model is 300 m long in the X direc-tion, 100 m in the Y direction and 80 m in the Z direction. The sim-ulation depth is 300 m. We applied displacement and stressboundary constraints on the pre and post, left and right and bot-tom surface of the model. The initial vertical stress wasrz ¼ cH ¼ 6:675 MPa, and the horizontal stress rx ¼ ry ¼ rz. Themechanical parameters of the rock mass are shown in Table 1.
3.1. Development of regular stress distribution in rock surrounding aroadway
After the downhill structures were mined, the stress in the rocksurrounding the coal, was re-distributed and became concentratedtowards the left of downhill material and to the right of the down-hill return air. The largest vertical stress was 13 MPa, which oc-curred 16 m towards the left of the downhill material and 15 mto the right of downhill return air. The stress concentration factorwas 1.95.
When mining started in the Wulimiao Coal Mine, the advancedsupport stress caused by mining and the stress in the surroundingrock towards the right of the downhill return air downhill weresuperimposed. The maximum vertical stress, located 14 m to theright of the downhill return air, was 17 MPa. The stress concentra-tion factor increased from 1.95 to 2.55. The maximum vertical
a b
Fig. 2. Failure condition of three downhill structures. (a) ‘‘Tip peach type’’ da
Table 1Mechanical parameters of rock mass.
Lithology Density (kg/m3) Bulk modulus (GPa) Shear modu
stress on the downhill material, located 15 m to its left, was14 MPa, with the stress concentration factor increasing from 1.95to 2.10.
When mining started in the #3209 upper layered working face,the advanced support stress caused by mining and the stress in thesurrounding rock towards the left of the downhill material weresuperimposed. The maximum vertical stress was 17 MPa, 13 m to-wards its left. The stress concentration factor increased from 2.10to 2.55. Neither the position nor the maximum value of the verticalstress in the surrounding rock, towards the right of the downhillreturn air, changed. Because stresses were released towards theinternal roadway space, the value of the surrounding rock stresswas significantly lower than that of the rock stress. The distribu-tion of the stress contours of the three downhill structures isshown in Fig. 3. The centers of the three downhill structures are130, 146 and 164 m from the lateral axis.
3.2. Regular distribution of rock surrounding the plastic zone of aroadway
When the downhill structures were dug in the coal, the stressconcentrated in a specific range of the plastic zone, a ‘‘cruciform’’distribution appeared in the rock surrounding the roadway, caus-ing damage to the rock surrounding the three downhill structures.The range of plastic zone of the roof and floor of the downhill beltwas 4.5 m, much larger than the downhill material and air return,while the range of plastic zone of both slopes is about 7 m, almostthe same as that of the three downhill structures, as shown inFig. 4a.
When mining started in the Wulimiao Coal Mine, the range ofthe rock surrounding the plastic zone, representing a ‘‘Q’’ distribu-tion, expanded significantly. The range of rock surrounding theplastic zone of the roof and floor was about 4.5 m, almost the samefor each of the three downhill structures. Both slopes of the plasticzone of three downhill structures are very different. The plasticzone of the floor of the three structures extended to the bottomboundary of the stable rock, as shown in Fig. 4b.
When mining started in the #3209 upper layered working face,the range of the rock surrounding the plastic zone, presented a‘‘pion’’ distribution which further expanded. The plastic zone ofthe floor of the three downhill structures extended to the bottomboundary of the stable rock. Coal pillars between the three struc-
c
mage, (b) ‘‘crooked peach type’’ damage and (c) ‘‘flat top type’’ damage.
Fig. 3. Distribution of vertical stress contours of three downhill structures. (a) Period of roadway driving, (b) period of mining one side, and (c) period of mining both sides.
a b c
Fig. 4. Distribution of plastic zones of rock surrounding three downhill structures. (a) Period of roadway driving, (b) period of mining one side; and (c) period of mining bothsides.
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roadway names
Roof Floor Left slope Right slope
Fig. 5. Deformations in rock surrounding three downhill structures.
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tures were damaged by shear failure. The range of plastic zone ofthe roof of the three structures expanded to about 7 m, but thefloor and both slopes of the plastic zones of the three structuresdid not change significantly, compared with one side mining, asshown in Fig. 4c. Ranges of the rock surrounding the plastic zoneof the three downhill structures are shown in Table 2.
3.3. Regular deformation of rock surrounding a roadway
When the downhill structures were dug in the coal, the totaldeformation of the rock surrounding the three downhill structuresbecame comparatively smaller, i.e., less than 180 mm. However,the deformation of the downhill belt was larger than that of thedownhill material and air return. Because of mining in the Wuli-miao Coal Mine, the deformation of the rock surrounding the threedownhill structures dramatically increased, with the maximumclose to 1 m. The bending deformation of both slopes of the down-hill material tended towards the internal roadway. A ‘‘tip peachtype’’ damage appeared in the vault of the downhill material. Seri-ous deformation appeared in the floor and both slopes of thedownhill belt. Clear bending deformation appeared in the left slopeof the downhill air return. A ‘‘crooked peach type’’ damage ap-peared in the vault of the downhill material. As of now, the threedownhill structures can no longer be worked safely, hence the firstrevisions were implemented in the three downhill structures. Be-
Table 2Ranges of rock surrounding plastic zone of three downhill structures.
Rock surrounding plastic area Material downhill Belt do
Period ofroadwaydriving
Period of oneside mining
Period of bothsides mining
Periodroadwadriving
Roof plastic area (m) 0.8 4.5 7.0 3.5Floor plastic area (m) 1.2 4.5 4.5 4.0Left slope plastic area (m) 6.0 14.0 18.0 7.0Right slope plastic area (m) 7.0 7.0 7.0 8.0
cause of mining the #3209 working face of the upper layer, a‘‘crooked peach type’’ damage and intense floor heave occurredin the vault of the downhill material and air return. A ‘‘tip peachtype’’ damage and severe floor heave occurred in the vault of thedownhill belt. However, the deformation in surrounding rock ofboth mining sides is smaller than that of one side mining. The con-crete deformations in the rock surrounding the three downhillstructures are shown in Fig. 5.
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4. Failure analysis
(1) Effect of stress superposition among roadway groups. Becauseof the narrow coal pillars between adjacent downhill struc-tures, the large deformation of the coal pillar between thedownhill material and belt was absorbed by the space ofthe downhill belt, under high stress. It also absorbed thedeformation of the coal pillar between the downhill airreturn and belt which enlarged the range of the rock sur-rounding the plastic zone of belt the downhills belt afterdriving the roadway.
(2) Effect of mining. The advance support stress caused by miningand the stress on the rock surrounding the downhill struc-tures are superimposed, which caused the stress on the rocksurrounding the downhill structures to be redistributed, theplastic zone expanded further and the surrounding rockfailed. Therefore, effect of mining is the most important fac-tor causing serious damage to the rock surroundingroadways.
(3) Lack of stability. The roof of the downhill structures is theupper layered top coal, which easily caves and is difficultto manage. The lithology of the downhill structures roof ismudstone, which easily leads to floor heave. After a seriesof downhill ripping, flitching and footwalling, continuouscracks appeared in the surrounding rock, its condition wors-ening constantly.
(4) Deformation of surrounding rock. Deformation of the sur-rounding rock is difficult to control with the existing archsupport. This arch is a rigid and passive supporting body,characterized by poor contractible performance and cannotbe adapted to withstand the large deformation of the road-way, affected by multiple mining. The interaction relation-ship between the arch support body and the surroundingrock is poor. This arch can be easily destroyed with a smallnon-uniform load. This makes it difficult for the supportingbody to play its supporting role completely.
5. Stability control technology
Roadways affected by mining abutment pressure are character-ized by broken surrounding rock in a large loose circle, easily lead-ing to caving and rib spalling accidents, thus affecting coal mineproduction [5,6]. It is difficult to control this kind roadway withbolt or arch support. Grouting is a technique for strengtheningfractured rock surrounding roadways, which fills and consolidatesdestroyed or original fractured surfaces, improve the strength ofrock masses, provide for improved conditions for surrounding rock,participates in equalizing its stress distribution and forces the sur-rounding rock to play its role in bearing loads [7–15]. It is an effec-tive way to control fractured rock surrounding roadways underconditions of complex stresses with an organic combination of gro-uting reinforcement and bolt support. Given the maintenancecharacteristics of the downhill material, belt and air return, wedetermined stability control methods for rock surrounding thethree downhill structures.
Install grouting bolt
Roadway driving
Arching supporting Drilling
Seal hole
Config grout
Test supporting effec
Optimize support parameters
Fig. 6. Technological process of stability control
(1) Grouting reinforcement behind dressed stone arch walling. Gro-uting reinforcement with high-water rapid hardening material cantransform the strength condition of the original fracture surfaceinto the strength condition of the coal in the rock surroundingthe three structures and improve the mechanical properties ofdownhill fractured surrounding rock for improving a bolt and cableanchor foundation and its anchoring face.
(2) After finishing grouting reinforcement behind the arch wall-ing, the roof and slopes of the roadway were supported with exten-sive whorl-steel bolting, and a resin explosive roll. Together with ametal net and a steel ladder beam, vertical and lateral groups ofinterrelated bolts were made, which formed an integrated supportstructure. We also used a ladder beam support. The trend andtransverse of the bolt groups are interrelated and the anchoragezones of rock surrounding a roadway form a comprehensive sup-port structure, greatly enhancing the bearing capacity of the roofand both slopes of the roadway.
(3) Small diameter pre-stressed roof anchor. A small diameter pre-stressed anchor was installed in the roof to improve the support,reducing the load of both slopes from the roadway roof andimprove the stability of the surrounding rock down slope.
(4) Test of quality of support. We adjusted and optimized the tim-ing of the support parameters, according to our measurement data.The concrete technological process is shown in Fig. 6.
6. Field experiment
6.1. Roadway grouting reinforcement
Grouting reinforcement was applied to the three downhillstructures with high-water rapid hardening material, which con-solidated the fractured surrounding rock. Grouting reinforcementparameters are shown in Fig. 7.
The grouting pressure applied was 1.5 MPa, with a maximumless than 3.0 MPa. The grout diffusion radius was determined as2.0–3.0 m by laboratory experiments and field measurements.The grouting holes were arranged in a row every 2 m with 6 holesin every row. The top grouting holes of the roadway were arrangedin 45� angles with a vault, with a span of the bilateral groutingholes of the roadway of 1000 mm. The diameter of grouting holewas 42 mm with a depth of 2400 mm.
6.2. Combined bolt and cable support
Bolts and cables were used to reinforce the support after grout-ing reinforcement of the three downhill structures. The supportparameters are shown in Fig. 8.
(1) Bolt support. We used 2400 mm long whorl-steels with20 mm diameters. Spans and rows were both 800 mm. Thewhorl-steel bolts were anchored with explosive resin rolls.For every whorl-steel bolt we used two explosive rolls, oftypes CK2335 and Z2360 respectively. The metal net andsteel ladder beam made from Ø14 round steel were usedin the roof and in both slopes of the roadway.
Grouting
Anchor girder net supporting
Reinforced supporting with anchort
Stop grouting after reaching design pressure grout amount
technology of rock surrounding a roadway.
Φ
Φ
Φ
Φ
Φ
Φ
a b c
Fig. 8. Layout of roadway support. (a) Material downhill, (b) belt downhill, and (c) return air downhill.
a b c
Fig. 7. Grouting hole layout of three downhill structures. (a) Material downhill, (b) belt downhill, and (c) return air downhill.
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Fig. 9. Deformation curves of rock surrounding three downhill structures. (a) 1. Roof and floor relative deformation of material downhill; 2. Roof and floor relativedeformation of belt downhill; 3. Roof and floor relative deformation of return air downhill; 4. Roof and floor relative deformation velocity of material downhill; 5. Roof andfloor relative deformation velocity of belt downhill; 6. Roof and floor relative deformation velocity of return air downhill; 1-Both slopes relative deformation of materialdownhill; 2-Both slopes relative deformation of belt downhill; 3-Both slopes relative deformation of return air downhill; 4-Both slopes relative deformation velocity ofmaterial downhill; 5-Both slopes relative deformation velocity of belt downhill; 6-Both slopes relative deformation velocity of return air downhill.
Y. Yu et al. / Mining Science and Technology (China) 21 (2011) 301–306 305
(2) Cable support. We used 7300 mm long anchor cables with a17.8 mm diameter. Spans and rows were 2000 mm and2400 mm respectively. For every pre-stressed anchor cablewe used three explosive resin rolls, one of type CK2335and two Z2360 rolls.
6.3. Control effect
In order to investigate the effect of the control of the rock sur-rounding the three downhill structures reinforced by groutingand bolt support, we established stations within the downhillstructures to measure roof, floor and slopes convergence. Fig. 9shows that after grouting reinforcement and bolt support, the rocksurrounding the downhill structures gradually stabilized. After54 days, the roof, floor and slopes of the downhill material con-verged almost unchanged. The convergence of roof and floor wasless than 200 mm. The convergence of both slopes was less than160 mm. After 72 days, the roof, floor and both slopes of the down-hill belt converged again almost unchanged. The convergence of
roof and floor was less than 200 mm. The convergence of bothslopes was less than 160 mm. After 90 days, the roof, floor and bothslopes of the downhill air return converged almost unchanged. Theconvergence of the roof and floor was less than 300 mm. The con-vergence of both slopes was less than 250 mm. Given this situa-tion, the roadway could be used normally, obtaining favorabletechnical and economic benefits.
7. Conclusions
(1) The main reason for failure in the form of deformation ofrock surrounding a roadway under mining abutment pres-sure, is the high stress caused by its superposition betweenadvance support stress from mining both sides and the sur-rounding rock stress after driving the roadway. Secondly,stress superposition among roadway groups, failed integrityof rock surrounding roadways caused by frequent revisionsand unstable support structures with arch support are addi-tional and important reasons leading to roadway damage.
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(2) Stability control of roadways must be adapted to the defor-mation features of the surrounding rocks. The deformationof rocks surrounding roadways was effectively controlledwith grouting reinforcement behind a dressed stone archwalling with high-water rapid hardening material and com-bined support with bolts and cables. In the end, the rockssurrounding the roadway come to a steady state in complexstress conditions.
Acknowledgments
Financial support for this work, provided by the National Natu-ral Science Foundation of China (No. 50774077), the ResearchFoundation of the State Key Laboratory of Coal Resources and MineSafety (No. SKLCRSM08X04), the National Basic Research Programof China (No. 2007CB209401), the Foundation for the Author of Na-tional Excellent Doctoral Dissertation of China (No. 200760), theProgram for New Century Excellent Talents in University (No.NCET-06-0475) and the Science Foundation for Youth of China Uni-versity of Mining and Technology (No. 2008A002). The authors aregrateful to Professor Bai Jianbiao, Lecturers Wang Xiangyu and CuiHua, Graduate students Chen Ke, Zhang Liang, Ma Shuqi, WangMeng and Ma Chuanyue for their help in part of the work and tothe coal mine mentioned for providing an experimentalenvironment.
References
[1] Hou CJ, Guo LS, Gou PF. Bolt support of coal roadway. Xuzhou: China Universityof Mining and Technology Press; 1999 [in Chinese].
[2] Kang HP, Wang JH. Rock bolting theory and complete technology for coalroadways. Beijing: Coal Industry Press; 2007 [in Chinese].
[3] Chen YG, Lu SL. The control of roadway surrounding rock of Chinamine. Xuzhou: China University of Mining and Technology Press; 1994 [inChinese].
[4] Lu SL, Tang L, Yang XA. Anchor-hold and anchoring technology. Beijing: CoalIndustry Press; 1998 [in Chinese].
[5] Bai JB. Surrounding rock control of entry driven along goaf. Xuzhou: ChinaUniversity of Mining and Technology Press; 2006 [in Chinese].
[6] Zhou HQ. Supporting limitation and stability function principle in roadway andits applications. Xuzhou: China University of Mining and Technology Press;2006 [in Chinese].
[7] Zhang N. Delay grouting control theory of surrounding rock in roadway and itsapplications. Xuzhou: China University of Mining and Technology Press; 2004[in Chinese].
[8] Chang JC, Xie GX. Mechanical characteristics and stability control of rockroadway surrounding rock in deep mine. J China Coal Soc 2009;34(7):881–4[in Chinese].
[9] Qi HG, Guo XX, Yu SQ. Deformation mechanism of greatly cracked tunnel andsupport technique by grouting and bolting. J China Coal Soc2008;33(11):1224–9 [in Chinese].
[10] Bai JB, Hou CJ. Control principle of surrounding rocks in deep roadway and itsapplication. J China Univ Min Technol 2006;35(2):145–8 [in Chinese].
[11] Lu Y, Liu CY. Similarity simulation of bolt support in a coal roadway in atectonic stress field. Min Sci Technol 2010(5):718–22.
[12] Józef K, Janusz M. Empirical-analytical method for evaluating the pressuredistribution in the hard coal seams. Min Sci Technol 2009(5):556–62.
[13] Yang J, Sun XM, Wang SR. Study on deformation and failure laws of deeproadway and countermeasures in Jining Coal Mine NO2. Chin J Rock Mech Eng2009;28(11):2280–5 [in Chinese].
[14] Kang HP, Wang JH, Lin J. High pretensioned stress and intensive bolting systemand its application in deep roadways. J China Coal Soc 2007;23(12):1233–8 [inChinese].
[15] Jia P, Tang CA, Wang SH. Destroy mechanism of tunnel with stratified roof. JChina Coal Soc 2006;31(1):11–5 [in Chinese].
