Journal of Mining and Safety Engineering

Research on the grouting diffusion mechanism and its application of grouting reinforcement in deep roadway

LIU Quansheng, LU Chaobo, LIU Bin, LIU Xuewei

2014, 31(3): 333-339.

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For the deficiencies of common grouting technology in deep soft rock roadways, three-step grouting technology is proposed to reinforce the surrounding rock to get better results. The three-step grouting diffusion mechanism is explored in this paper, and the technique is applied to field engineering. Step one is to grout with low grouting pressure to fill the gravel pack interval and blocked the channels which relieves the grouting pressure in the shallow surrounding rock. Step two is mainly for the filling of the fractures space in the surrounding fractured rock to improve the bearing capacity of the surrounding rock through the effect of grout cementation and filling with grout solidifying. In step three grouting is prepared to further reinforce the strong deformation sections of the roadway. Although increasing the working procedures of grouting, the three-step grouting technology ensures the radius of grouting diffusion. The inspection holes monitoring shows that grouting diffusion depth is generally more than 3 m, and that the three-step grouting technology has achieved the expected effective range of grouting reinforcement. The roadway deformation monitoring datum of grouting before and after indicates that three-step grouting technology has greatly improved the effect of grouting reinforcement of surrounding rock and effectively improved the bearing capacity of rock mass，and thus the overall stability of surrounding rock of roadway is effectively controlled.

Study on asymmetric distortion and failure characteristics and stability control of soft rock roadway

YU Yang, BAI Jianbiao, WANG Xiangyu, SHEN Wenlong, LIAN Changjun

2014, 31(3): 340-346.

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In response to the issues of asymmetric distortion and failure of roadway and hard maintenance after excavation of support in Meihe coal mine，field geological investigation，rock mechanics experiment，microstructure materialization experiment，theoretical analysis and numerical simulation calculation are used synthetically in this paper to analyze the main factors responsible for asymmetric distortion and failure of the roadway. The research found out that the process of the roadway distortion and failure has significant stages and revealed the distribution of stress and plastic zone in the condition of non-support，the asymmetric distortion and failure characteristics of roadway under the condition of original symmetric support. On this basis, the paper proposed the concept of ‘weak structure’ and asymmetric control technology of ‘roadway overall stability，weak structure reinforced support’，and industrial test is implemented in the field. The practice proves that the compatible deformation of support structure and surrounding rock is realized，the deformation of surrounding rock of roadway is controlled remarkably and avoided getting into the status of long-term rheological; the overall stability of roadway is guaranteed by the asymmetric control technology.

Stress control of deep cutting along roadway over roof rock

LIU Zhenghe, YANG Lusheng, SONG Xuanmin, ZHAO Yangsheng, FENG Zengchao, YANG Dong

2014, 31(3): 347-353.

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When there is a layer or several layers of hard rock seam roof, the stress concentration in lateral rock increases with increase of mining height, which leads to the widening of the roadway coal pillars. In order to reduce the width of the roadway coal pillars, a technological principle of cutting different depth fracture along roadway in roof is proposed. A systematic study of the surrounding rock stress transfer and control is conducted through similar simulation and numerical simulation when roadside roof has been cut at different depth. The impact of cutting fracture depth on strata breakage and roof sinking is revealed. The results show that: with the increase of kerf depth, the rock stress of goaf decreases gradually; the higher the rock strata is, the smaller the stress is. With the increase of cutting fracture height of the lateral goaf roof, the rock stress above the coal pillars increases gradually; the higher the rock strata is, the bigger the stress becomes. The rock stress above the coal pillars(10 m and 20 m) and the distance from peak stress to cutting fracture edge are in nonlinear inverse relationship with cutting depth. While the strata stress at 10 m and 20 m above the goaf shows an exponential relationship with cutting depth. It means that the deep cutting can effectively control the distribution of strata stress, stress maximum and the distance of peak point from the edge of cutting fracture.

Ultimate self-stable arch theory in roadway support

HUANG Qingxiang, LIU Yuwei

2014, 31(3): 354-358.

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Based on the homeostatic phenomenon in roof caving process of roadway, the concept of ultimate self-stable arch of roadway is put forward. In accordance with the principles of roof rock destruction pull caving area and no tensile stress conditions, the limits of homeostasis arch elliptic curve equation is set up. Considering the impact of roadway ribs damaged areas on the roof stability, the ultimate self-stable arch theory in roadway control is proposed. Based on the interaction of “floor-rib-roof”, the roadway support should be designed according to the whole ultimate self-stable circle. Combining plenty of practical work experiences, the roadway support principles are renewed to pay attention to the ribs and floor support, and the key points of bolting quality are presented. This research will provide a new scientific basis for the theory and practice of roadway support.

Combined test research on coal pillar width setting of district sublevel for fully-mechanized face with large mining height based on 3D dynamic strain monitoring

ZHENG Yangfa, JU Wenjun, KANG Hongpu, WANG Zhanling, JIANG Pengfei

2014, 31(3): 359-365.

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Based on the theoretical calculation of the ideal elastic-plastic mechanics, FLAC3D numerical simulation analysis and multiple mining engineering practice, it could be judged that the forty-meter width of the coal pillar is significantly larger in Ningdong test mining sublevel. The in-situ underground combined test is implemented to have a real-time monitor of sixteen hollow inclusions’ 3D strain, the dynamic change of stress and deformation of the surrounding rocks in test roadways with coal pillars of different width (ten meters and twenty meters) pre and post the mining-induced influence of the working face. Test results show that there exists obvious core zone with significantly lower memory elastic stress in the coal pillar of forty-meter width, which is a favorable location for the roadway layout; the ad- vanced influence radius of mining-induced pressure is about forty-five meters, with the advance of the work face, mining-induced stress peak is gradually passed to the depths of the coal pillar during which stress increment constantly attenuates. Affected by mining-induced function, the original equilibrium state of the coal pillar 3D principle stresses gradually deteriorates, and, the test roadway with coal pillar of 10-meter width suffers less deformation than that of 20-meter width. Based on the above theoretical calculations and underground test results a reasonable width scope of district sublevel coal pillar design can be found (8.9 m to 12.5 m) in the test mine under specific geological condition, which provides empirical guides for the width design of district sublevel coal-rock pillar under similar mining conditions.

On the performance of lengthened bolt coupling support system in roadway roof

LIU Hongtao, WANG Fei, JIANG Lishuai, ZHAO Xidong, WANG Guanghui

2014, 31(3): 366-372.

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In view of the force-deformation incoordination problems of traditional anchor and cable in largely deformed roadway roof support system, a new type of lengthened bolt is invented. After a tensile test and mechanical analysis of traditional anchor, cable and new lengthened bolt and an analysis of the deformation and failure characteristics of surrounding rock in large deformation roadway roof, ,a support system for large deformation roadway roof called lengthened bolt coupling support system is devised. The constitutive models of lengthened bolt coupling support system and cable coupling support system are established and analyzed. The research result shows that when the roof deformation exceeds the maximum elongation of cable, the cable in the coupling support system reaches the elongation equilibrium and is broken earlier than the traditional bolt due to the low elongation, but lengthened and traditional bolt could coordinately support the surrounding rock of roof both in deformation and force aspects. Lengthened bolt could improve the self-loading performances of surrounding rock, and not only has a large amount of elongation, but ensures a strong supporting force. Lengthened bolt coupling support system has been successfully applied to the test area of 11031 in Zhaogu Mine. The deformation of new support program is 20% less than that of the ordinary program. Besides, time-cost decreases 33% from roadway excavation to a stable roof rock, and the subsidence speed turns to be much slower than that of an ordinary program.

On the deformation and failure characteristics of the Tertiary soft rock roadway and coupling control measures

YANG Jun, YU Shibo, TAO Zhigang, SUN Xiaoming, WANG Dong

2014, 31(3): 373-378.

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Based on the analysis of the deformation and failure characteristics of the Tertiary soft rock roadway in China, the further research on the successful supporting practice of bolt-net-cable-double- layer truss in Liuhai coal mine has been carried out . The result of the field investigation shows that the tertiary soft rock roadway has been damaged in wide range with high repair rate, complex damage forms and nonlinear large deformation．The bolt-net-cable-double-layer truss support technology is put forward based on the comprehensive study of the compound deformation mechanism of high stress, jointing and intense swelling soft rock. Because of the characteristic of the coupling stress, the roadway deformation can be controlled effectively in time and space with the aid of bolt-net-cable, and it is also possible for the deformation and stress in the surrounding rock to become uniform. The anchor cable can take full advantage of the strength of deep hard surrounding rock, while the double layer truss can bear the rheological load of soft rock with high strength, wholeness, cooperativity and uniform contacted stress, making possible the effective transformation of high stress in the deep surrounding rock of roadway. Engineering practice shows that the new technology has been successfully applied to ensure the stability of roadway.

Study on buckling instability of surrounding rock based on cusp catastrophe model

LI Ming, MAO Xianbiao, MAO Rongrong, TAO Jing

2014, 31(3): 379-384.

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Based on the stability characteristics of surrounding rock of deep roadway, the forming mechanism and stable mechanics model of crack plate structure are studied. With the application of cusp catastrophe model on nonlinear stable structure, the instability property of crack plate structure is analysized and the necessary and sufficient conditions under two different cases is put forward. The study proves that the catastrophe theory can effectively predict the dangerousness of slough roadway rock burst. By reinforcing surrounding rock on the top and bottom corner, the stiffness of surrounding rock is increased and its anti-buckling capacity is also enhanced, therefore rock-burst accident in spalling- type coal mine tunnel can be effectively prevented.

Chaotic prediction of surrounding rock deformation speed in coal roadway supported by bolt

LIU Xuesheng, TAN Yunliang, NING Jianguo

2014, 31(3): 385-389.

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Bolting surrounding rocks of coal roadway have nonlinear characteristics. In this paper, based on the largest Lyapunov exponent of deformation speed series, progressive forecasting and non-progressive forecasting methods of surrounding rock deformation speed were established to forecast the horizontal approaching speeds in two sides of one roadway in Muchengjian coal mine. The results show that the deformation speeds of surrounding rocks have chaotic characteristics, and the chaotic degree of deformation speeds between two sides is higher than that between the roof and floor. The fitting effect of surrounding rock deformation speed is satisfactory, and the fitting error is small. The prediction effect of progressive forecasting is better than that of non-progressive forecasting, and the largest predictable time is about 45 days, with the average relative error of 10.99%. Based on the progressive forecasting results, supplement support practice of surrounding rocks was carried out, and a satisfactory result has been achieved.

On danger-area forecast system of working face opening roof caving

LI Ji, FAN Long, ZHAO Xidong, WANG Fei, YANG Guangrong

2014, 31(3): 390-398.

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In order to improve the precision and accuracy in forecasting the risk level of opening roof only applying single index, affecting factors and mechanism of roof caving in opening area are studied with methods of RFPA numerical simulation and analysis hierarchy process. The mechanism shows that the difference in rock strength and stiffness leads to a dramatic variation in the separation size between roof strata, which has a direct impact on the degree of roof carving. Also, it points out that the anchor bolt (cable) supporting quality and the structure of roof strata are two main factors affecting roof caving in opening area. Taking these two factors as forecast indexes, the danger-area forecast system of opening roof was built up, and forecast of the danger-area is made in opening area of 12018 working face basing on this system. The forecast result illustrates that roof caving danger exists in more than 50% area of the roof in opening area, and the danger-areas distribute in the upper and lower end position and in the central position. According to this forecast result, reinforcing supported the danger-area to ensure the safety and reliability of mining operation, with a separate minor average size of 12 mm and a gradual falling rate of 0.2 mm/d.

A simple calculation method for roof first caving span based on the plate structure theory

WU Fengfeng, LIU Changyou, YANG Jingxuan

2014, 31(3): 399-405.

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To analyze the roof first caving span based on the plate structure theory, considering the roof strata are loaded by continuous linear action, the roof first caving spans in the simply supported and clamped boundaries were investigated respectively by using the theoretical analysis and field measurement. In this paper, the relationship between the bending moment in the rupture peril point of roof and ratio of roof length in the two directions (strike and inclined directions) were obtained, and the largest bending moment expression under the random length ratio of the two directions were provided. Meanwhile, the largest stress distribution expressions in the simply supported and clamped roof boundaries were obtained when considering the trapezoidal loading of the inclined roof. Moreover, a simple calculation method for roof first caving span based on the plate structure theory was proposed. According to the contrast results between the field measurement and theoretical analysis of the roof first caving span in 8281 coal face in Taoyuan mine, Huaibei mining area, the measured first caving spans which are ranging 25~34 m basically match the theoretical calculation results (26.5~35.5 m), which indicates that the calculation method has good adaptability for the roof first caving span forecasting.

Stability of surrounding rock in head face of upward fully-mechanized caving face and its control technology

GUO Weibin, LU Yan, HUANG Fuchang, LIU Changyou, BUI Manhtung, DO Anhson

2014, 31(3): 406-412.

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In this paper, to ensure the stability of the coal wall and head face roof, and realize the safe and high efficient production in the upward full-mechanized caving face, the instability characteristics of coal wall and immediate roof was analyzed, the mechanics model and condition to keep the stability of coal wall and head face roof were established, and the interaction between the setting load of support, the front angle of column, and the upward mining angle was discussed. The results indicate that the increasing of upward minging angle, the improving of setting load and the reducion of the front angle of column, are beneficial to control the stability of the coal wall and head face roof. When the front mining angle is less than 11°, the key point is to control the roof subsidence, otherwise, the key point is to control the rib spalling and roof downslide in the face. Combined with the field conditions of No.80113 up- ward full-mechanized caving face in one mine, the reasonable setting load of support and front angle of column were defined, and the pertinence measure to control the stability of surrounding rock was put forward, which has ensured the safe and high efficient production.

The law of rock pressure in the stope with blocking mining by the continuous miner

ZHOU Maopu, CAO Shenggen, JIANG Xiaojun

2014, 31(3): 413-417.

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With the corner coal mining conditions in Wulanmulun Coalmine of Shendong Company as the engineering background, mining pressure characteristics under technological conditions of blocking mining by the continuous miner is researched numerically through the software of FLAC3D finite difference combined with on-the-spot pressure observation. The research finds that the use of complete management of roof caving method will, as the working face advances, make the gob area expand unceasingly, and thus the peak stress and stress of coal pillar between knife caving change drastically, with the stress concentration factor ranging from 2.2~2.7, and that the maximum vertical displacement position of the direct roof in the mined-out area will extend towards mining, which causes abscission layer of the direct roof off the main roof. The abscission layer becomes larger and larger and will collapse extensively in the triangular area formed by the headway and the roadway.

Physical simulation on strata behavior of large mining height fully mechanized face in shallow-buried and thick seam

LAI Xingping, DAN Pengfei, ZHENG Jianwei, CAO Jiantao, CUI Feng, WANG Chunlong

2014, 31(3): 418-423.

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According to conditions of No.5-2 super-thick coal seam in Zhangjiamao coal district of Yushen mining area, the physical and mechanical properties of coal-rock masses were measured by lab mechanics experiments. The physical simulation model was built up, and the mechanical properties of support were simulated．The simulation results indicate that the strata behavior of large mining height fully mechnized face is dramatic．Meanwhile, the deformation and damage of rock masses with the excavation were monitored by the acoustic emission (AE) and support pressure monitoring, and the strata movement and evolution of abutment pressure during roof weighting were analyzed．The results also show that the wave variation from AE is remarkable with abutment pressure change, and the maximum events are reached to 2578 per min．The step of periodical weighing is about 15.0 meters, the advanced abutment pressure zones range from 10.0 to 13.0 meters ahead of the workingface, and the peak pressure moves toward the direction of mined-out area．Moreover, the support with the rated working resistance of 12 000 kN can satisfy the efficient and safe excavation of the coal seam.

Experimental research on realizing inclined successive and sufficient pressure-relief in pillarless protective coal seam mining

WANG Zhiqiang, ZHOU Lilin, YUE Yucheng, ZHANG Lihai, ZHANG Zheng, ZHAO Kecheng, ZHAO Jingli

2014, 31(3): 424-429.

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The research in this paper was conducted based on realizing inclined sufficient and successive pressure-relief in lower protected coal seam mining. The results show that due to the length of coal face, interval between layers, borderline of pressure-relief and coal pillar layout, the pressure-relief in the inclination of the distant protected coal seam can’t be realized, and the high stress zones or insufficient pressure-relief zones may exist in the close protected seam. The safe and successive mining of protected seam can’t be realized. On this basis, the pillarless face layout in protected seam was proposed. Thus, the overlying strata movement during the successive face mining may integrate with that of the first mining face, the inclined successive mining length can be enlarged, and the mining degree of the overburden strata is more sufficient. So, the horizon of the protected seam in the fractured zone is relatively lower, which means the pressure- relief degree is more sufficient. Hence the inclined successive and complete pressure-relief in the protected coal seam can be realized. Combined with the practical engineering conditions in one mine, taking recovery ratio and engineering amount of roadway into consideration, the profit of No.2 Coal Seam mining when adopting pillarless mining is increased by 284 million yuan compared with the original mining plan. Morevoer, compared with the roadway driving along the goaf remaining 5-metre pillar, the engineering expense of rock roadway can be reduced by 6 million, and the mining profit can be increased by 140 million.

Space-time rule of the control action of filling body for the movement of surrounding rock in method of the delayed filling open stoping

YU Shibo, YANG Xiaocong, DONG Kaicheng, XIE Lianku, SUN Xiaoming, GUO Lijie

2014, 31(3): 430-434.

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Selecting the first mining project at the -230~-290 m middle section of an iron mine as the research object, the space-time rule of filling body for controlling the movement of stope surrounding rock in the large-diameter long-hole with the delayed filling open stoping method has been studied in methods of field monitoring of multipoint displacement extensometers and numerical simulation. The three-dimensional mechanical structure model of abutment pressure for the movement control of overlying rock mass has been established in this paper. The results indicate that the movement and deformation of surrounding rock could be effectively controlled as long as the stope of 15-meter width has been filled within the demurrage time. The footwall becomes the main bearing structure to support overlying rock mass. The capacity of the cement filling body is limited to support overlying rock mass while it is a valid constriction of the movement and deformation of overlying rock mass. The maximum vertical deformation of overlying rock mass is 39 mm and there is almost no difference between the ultimate deformations of surrounding rock at the room and pillar. So the ground pressure is not obvious.

Influence of blasting vibration and structural plane progressive failure on slope stability

REN Yuelong, CAI Qingxiang, SHU Jisen, ZHOU Wei, HAN Liu

2014, 31(3): 435-440.

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In order to investigate the impact of blasting vibration and structural surface damage on slope stability and the stability coefficient variation and undermining mechanism during blasting process, the research gets the law of the shear strength variation during under mining process with the degradation mechanism for shear strength in structural surface as the basis. In combination with the mechanism of blasting dynamic loading, calculation formula of both sliding and sliding resistance forces under the condition of fixed vibration were corrected. The calculation methods of timeliness stability coefficient for planar and folding landslide models are deduced and applied to the study of the north slope stability at Haerwusu Surface Coal Mine. The results show that stability factors are 1.348 and 1.173 for intact structural surface and damaged structural surface respectively. In addition, the slope stability coefficient decreases linearly during the slope undermining progress while the decreasing speed is proportional to the cohesive force C. Furthermore, under the action of blasting vibration, the slope timeliness stability fluctuates up and down in response to seismic acceleration and the combined influence of both blasting vibration and structural surface damage makes the slope timeliness stability present a descending trend when fluctuating.

Modeling experimental method of unsteady heat conduction in surrounding rock of roadways or tunnels

ZHANG Yuan, WAN Zhijun, ZHOU Changbing, CHENG Jingyi

2014, 31(3): 441-446.

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Based on the analysis of the thermal conduction characteristics of the surrounding rock of roadways or tunnels, the temperature field of the surrounding rock of roadways or tunnels is simplified as the unsteady heat conduction problems of a one-and a half dimensional infinite hollow circular sheet of insulation on both sides; the mathematical model for the temperature field is established under cylindrical coordinate, and the solution conditions related are explored; using equation analysis method and dimensional method, the paper deduces that the conduction similarity criterions of the surrounding rock of roadways or tunnels are the Fo, the Bi, the R in the conduction process, and the Nu, the Re, the Pr in the period of heat convection between the surrounding rock and the aircurrent. According to the similarity criterion, the paper indetifies that there are similarities between the the thermal conductivity prototype and model of the surrounding rock of roadways or tunnels, and certifies the feasibility of modeling experiment method with an example of drivage roadway.

Measurement analysis of overlying strata movement and surface subsidence by UCG strip mining

XIN Lin, WANG Zuotang, HUANG Wengang, ZHAO Kexiao, JU Yuanjiang, HE Sheng

2014, 31(3): 447-455.

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To obtain the surface sinking regularity in the case of super-subcritical extraction of shaft pillar by underground coal gasification (UCG) at former Ankou Coal Mine in Huating, Gansu Province, the first observation station in China for overlying strata movement and surface subsidence of strip-partial UCG working face is established. After the analysis of dynamic subsidence progress of surface measuring point and magnetic rings in surface shallow base point observation drilling, it is found that the subsidence curve shows the fluctuant and stepwise subsidence pattern which has five stages (sinking → rising → sinking → rising → sinking), and the surface movement and deformation of super-subcritical extraction of strip-partial UCG working face are obviously characterized by nonlinearity and discontinuity. The strata separation area is positioned by analyzing the dynamic subsidence progress of different magnetic rings. The geophysical prospecting finds that the projection plane of combustion space area seems like an oval, wide in the middle and narrow at both ends. The greatest subsidence angle, boundary angle of incline main section of subsidence basin and the maximum surface movement and deformation parameters (imax = 1.385 mm/m, εmax = 0.516 mm/m, Kmax = 0.275 mm/m2) are obtained, which indicates that the subsidence basin doesn’t reach the critical boundary, and this extraction of strip-partial UCG working face doesn’t form apparent subsidence basin and destroy the surface building in the observation period.

Analysis of deformation and strength characteristics of coal samples under the triaxial cyclic loading and unloading stress path

SU Chengdong, XIONG Zuqiang, ZHAI Xinxian, GU Ming

2014, 31(3): 456-461.

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The triaxial cyclic loading and unloading is conducted on the coal sample through RMT-150B rock mechanics testing system. The deformation and strength characteristics of the coal samples during the tests are analyzed. The results show that during the tri-axial cyclic loading the coal sample deformation shows distinct memory and good fitting in stress-strain curve between the cyclic loading and the continuous loading. For loading and unloading prior to yield, the elastic modulus of loading procedure is always lower than the elastic modulus of unloading procedure. And the elastic modulus increased slightly with the frequency. The elastic modulus of unloading procedure after peak value is lower than that before peak value; however, the elastic modulus is still higher than that of first loading procedure. The peak strength and residual strength are proportional to the confining pressure during triaxial compressive test, which is consistent with the Coulomb strength criterion. The friction factor obtained from the regression analyses of the peak strength, residual strength and confining pressure are roughly equal and the cohesion is reduced by 54.4%.

An experimental study of the influence of seepage pressure and initial porosity on variable mass seepage for broken mudstone

WANG Luzhen, CHEN Zhanqing, KONG Hailing, NI Xiaoyan

2014, 31(3): 462-468.

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In order to study the water inrush mechanism of Karst collapse column in different depths with coupled underground water pressure and compaction degree, a modified variable mass seepage experiment system on broken rock was used to conduct variable mass seepage experiments for broken mudstone considering particle migration. The duration time of seepage burst, weight of gushed particle, change rate of gushed particle weight and change rate of porosity were recorded and calculated, and the change rules under seepage pressure and initial porosity were obtained. It shows that in the initial seepage stage, seepage pressure and flow are stable, they change rapidly when seepage burst happens, and the change range is related to weight of gushing particle and the process of pore structure adjusting. With the initial porosity and seepage pressure increasing, the duration time of seepage burst gets shorter, weight of gushed particle gets larger, and these two parameters are fitted by logarithmic functions relating to seepage pressure and initial porosity; change rate of gushed particle weight and change rate of porosity get faster and both are fitted by exponential functions relating to seepage pressure and initial porosity. With seepage pressure increasing, the influence of initial porosity on the duration time of seepage bursts, weight of gushed particle, change rate of gushed particle weight and change rate of porosity are gradually weakened.

Study on coalmine boreholes shrinkage rule crossing swelling rock under the humidity stress field

LU Yiyu, HOU Jifeng, YOU Yi, GE Zhaolong, ZHANG Lei, AO Xiang

2014, 31(3): 469-475.

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In response to the issue of boreholes shrinkage crossing swelling rock in coal mine, in this paper，based on the humidity stress field theory, analytical solution of boreholes displacement considering humidity damage and expansion has been deduced by using finite element analysis program and introducing humidity corrected coefficient ψ(w).So，the effects of the humidity field and stress field on boreholes shrinkage are systematically studied and analyzed by the software ANSYS and similar model experiment. The results show that with the water ratio of surrounding rock increasing, the wall displacement of boreholes is nonlinearly growing; the wall is“dumbbell-shaped”with two ends crude and intermediate fine; the maximum radial displacement is in the central wall of the boreholes；the effect of lithology softening resulting from the boreholes shrinkage is obvious；as ground stress increases, the wall displacement of borehole increases linearly but has a small increase；the above two fields have enormous effects on boreholes shrinkage. As far as the degree of influence is concerned， humidity field causes the larger influence, followed by stress field.

Influence of pipe length on coupled relation between flow velocity of detonation front and overpressure

LIU Qian, LIN Baiquan, ZHU Chuanjie, JIANG Bingyou, HONG Yidu, SUN Yumin

2014, 31(3): 476-482.

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To study the influence of pipe length on the coupled relationship between the flow velocity of detonation front and peak overpressure caused by gas deflagration, the propagation characteristics of overpressure and velocity in a closed end pipe was studied by numerical simulation in this paper. The results show that all of the last explosion overpressure attenuate to 0.7~0.9 MPa in the pipes with different lengths. For the short pipes, the peak value of velocity is relatively small, and the decline trend of the peak velocity value is relatively gentle, while the oscillation frequency of velocity is relatively fast in the positive and negative interval. For the relatively long pipes, the peak value of velocity is relatively large, and the decline of peak value is fast, while the oscillation frequency is small. In addition, there exists a critical length between 15 m and 20 m. When the pipe length is larger than the critical length, two minimum overpressure peaks will appear, and when the pipe length is shorter than the critical length, only one minimum overpressure peak appears. Moreover, the first peak value of velocity is in good proportion to the overpressure. The coupled relation between the first peak value of velocity and overpressure was obtained by the numerical data fitting. The results can provide reference to explosion prevention and control in the limited space.

Partial W type ventilation system and its parameters optimization of longwall mining with sublevel caving

LI Yingming, XU Jicheng, ZHANG Han, YANG Mingdong, FU Yonggang

2014, 31(3): 483-488.

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In view of methane overrun with U type ventilation, Partial W type ventilation system is put forward. Gas emission law is studied by numerical simulation. Working face parameter is optimized to improve extraction rate as much as possible on the premise of mining safety. Research results show that gas concentration of working face, return current and upper corner drop sharply and methane overrun problem with U type ventilation is solved effectively; when coal pillar is 15 m wide, it will be stable in the course of drifting and mining. Field application of above findings ensures safety and high efficiency mining. The ventilation system provides a new method to realize the safe and efficient mining of high gas longwall mining with sublevel caving.

Influence of vacuum chamber volume on gas explosion suppression

SHAO Hao, JIANG Shuguang, LI Qinhua, WU Zhengyan, ZHANG Weiqing, WANG Kai

2014, 31(3): 489-493.

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In this paper, to study the correlation between vacuum chamber volume and its suppression of gas explosion, a L-type pipe of gas explosion was designed, and the round plastic plate which can adjust the volume of vacuum chamber was made. In addition, a series of experiments about gas explosion suppression were carried out with the vacuum chamber volume changed in descending order. The results show that the suppression effect of gas explosion depends on the relation of vacuum chamber volume and its critical volume. When the vacuum chamber volume is greater than its critical volume, vacuum chamber has good suppression effect of gas explosion. When the vacuum chamber volume is smaller than its critical volume, the vacuum chamber has no suppression effect, furthermore, the overpressure of gas explosion is higher and flame signal strength of gas explosion is stronger in comparison with no vacuum chamber. Meanwhile, for the pipe used in these experiments, the critical volume of vacuum chamber is 0.026 m3, which is 43% of the pipe volume. This conclusion displays that when the vacuum chamber volume is greater than 43% of the experimental pipe, the vacuum chamber has the suppression effect of gas explosion. Conversely, there is no suppression effect of gas explosion.

Action mechanism of methane gas in the process of coal blasting damage and fracture

CHU Huaibao, WANG Jinxing, YANG Xiaolin, YU Yongqiang, LIANG Weimin

2014, 31(3): 494-498.

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In this paper, the action mechanism of methane gas in the process of coal blasting damage and fracture were analyzed comprehensively based on the experiment, numerical simulation, and theoretical analysis results. The process of coal blasting damage and fracture includes the initial stage acted by blasting wave and the later stage acted by detonation gas and methane gas. The methane gas has positive role in the whole process, which can increase the blasting damage degree and the crack propagation velocity. In the initial stage, the methane gas can increase the stress wave peak, extend the action time and reduce the decay rate of stress in the tensile phrase. In the later stage, the methane gas takes pate in the crack propagation process drived by detonation gas, and the crack propagation will further promote under the superposition action of methane gas pressure field and detonation gas quasi-static stress field.

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