Journal of Mining and Safety Engineering

Failure mechanism and grouting reinforcement technique of large mining height coal wall in thick coal seam with dirt band during topple mining

WANG Jiachen, YANG Yinchao, KONG Dezhong, PAN Weidong

2014, 31(6): 831-837.

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In this paper, the general mechanism of rib spalling was analyzed in detail, the main failure form of coal wall was summarized combined with the real conditions of 2612 coal face in Dongpang Mine, and the mechanical model of rib spalling in topple mining face with large mining height was established. The results show that the roof pressure, dirt band thickness, topple mining angle, coal cohesion, et al, have a great influence on the coal wall stability, and relieving the roof pressure of coal wall is the most effective measure to prevent coal wall failure. Meanwhile, the reasonable drilling position and grouting reinforcement parameters of coal seam with dirt band were studied by theoretical analysis, UDEC numerical simulation, etc. The studies show that the grouting holes drilled in the breakdown point of rib spalling is the best solution. In addition, the coal wall grouting process and implementation effect in 2612 coal face with large mining height was introduced.

Overburden structure influence to support crushing and water inrush during mining under unconsolidated confined aquifer

WANG Xiaozhen, XU Jialin, ZHU Weibing, HAO Hao

2014, 31(6): 838-844.

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In this paper, in view of the support crushing and water inrush disaster when mining under unconsolidated confined aquifer, numerical and physical simulations were used to study the influence of overburden thickness, key stratum location and structural combination characteristics, etc., on support crushing and water inrush disaster. The results show that under the remarkable load transfer faction of unconsolidated confined aquifer, support crushing and water inrush disaster is mainly related to the overburden structure. The thinner the thickness of bedrock is, and the thicker the first key stratum above coal seam is, the more easily the integral breakage of overburden occurs if the super-thick stratum exits within 10 times mining height distance above coal seam, which may easily induce support crushing and water inrush disaster in coal face. Above has been validated at No.7114 and No.7112 working faces in Qidong Coal Mine. Based on the overburden structure, the dangerous area of support crushing and water inrush can be predicted before coal mining, which may provide a basis for such disaster prevention.

The theoretical research on basic characteristics of backfilling hydraulic support and its application

ZHANG Qiang, ZHANG Jixiong, TAI Yang, JU Feng, SUN Qiang

2014, 31(6): 845-851.

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Backfilling hydraulic support is key equipment which not only has the function of supplying the space for coal mining as traditional hydraulic support, but also has the special function of supplying the space for solid materials backfilling. The concept of backfilling characteristics has been approved because of the special function, meanwhile, the basic implication of backfilling characteristics has also been analyzed through the following six aspects: supporting intensity, supporting intensity of rear roof canopy, movement feature, compacting force, roof gap and the geological conditions adjustment feature. The backfilling characteristics is mainly influenced by support structure, geological conditions, backfilling technique, during which the support structure plays an important role; besides, thickness, roof and floor conditions, dip angle and depth of coal seam also have a huge influence on the effects of backfilling characteristics; and the backfilling technique will influence the backfilling performance by changing the value of roof gap and compacting force and the adjustment feature to different geological conditions. The effect of backfilling characteristics can be improved through working face layout optimization, structure design optimization, backfilling technique optimization and real time online monitoring during engineering practice. The research result has been proved by the using of ZZC10000/20/40 backfilling hydraulic support with six props and four-bar linkage in Jining No.3 Coal Mine.

Experimental study on engineering characteristics of super-high water filling body

WANG Xufeng, SUN Chundong, ZHANG Dongsheng, LI Yongyuan, XU Mengtang, GUAN Kai

2014, 31(6): 852-856.

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According to the space position between the strata in the gob caving and filling slurry, super-high water material god filling cementing body is divided into four basic forms, namely pure super-high water material cementation body(class 1), super-high water materials and waste rock mixed materials in lower half and pure super-high water material in upper half (class 2), super-high water materials and waste rock mixed materials in upper half and pure super-high water materials in lower half (class 3), and the super-high water materials and waste rock overall mixed materials (class 4). The uniaxial compressive strength, tensile strength and shear strength mechanics parameters have been tested, and a mechanical model of “filling body-main roof” for backfilling mining has been established. Due to the mechanical model, the influence of the material parameters on limit span of the main roof has been analyzed. Study results show that, during the process of pressing, the features of stress-strain of the four types materials are nearly the same, and the class 4 material has the largest compressive strength, higher than class 1 by about 20%. The tensile strength and the shear strength of class 2, 3, 4 are commonly higher than class 1, and the maximum value can be increased by 40% and 50%. The analysis results of the influence of the material parameters on limit span of the main roof show that, the broken immediate roof mixed into the pure super-high water material and formed the materials of class 2, class 3 and class 4 can increase strength of the filling body, which are of benefit to the stability of the main roof for backfill mining.

The research and application of longwall-roadway cemented backfilling mining technology in extra-thick coal seam

DENG Xuejie, ZHANG Jixiong, ZHOU Nan, AN Tailong, GUO Shuai

2014, 31(6): 857-862.

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In this paper, the longwall-roadway cemented backfilling mining technology is proposed to solve the engineering technical problems in extra-thick coal seam, such as low recovery rate and the difficulty in controlling the mining surface subsidence and overlying strata movement and deformation. Systematically, its layout of filling mining system, main equipment and technics are explained based on studying the principle of longwall-roadway cemented backfilling mining technology in extra thick coal seam. Moreover, the composition and bearing characteristics of the backfilling materials are studied to determine the optimal ratio of cemented filling materials which is well applied to field work, namely the optimal proportion of fly ash, lime slag, cement and gangue are 35%, 10%, 2% and 53% respectively. Furthermore, the slurry concentration is 79% and its uniaxial compressive strength can reach 1.8 MPa. The field engineering application shows that the filling ratio can reach 96.1% and the maximum surface subsidence value is 265 mm when mining 21 meter thick coal seam, namely the surface sinking coefficient is only 0.012 6, which shows an excellent application effects.

Rock burst danger monitoring based on microseismic method

YANG Chundong, GONG Siyuan, MA Xiaoping, WANG Guifeng

2014, 31(6): 863-868.

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Massive mining-induced seismic events may appear before the occurence of strong tremor, which is important information for the evaluation of rock burst danger. In this paper, based on the seismic monitoring by SOS Seismological Observation System installed in No.10 Coal Mine of Pingdingshan Mining Area, the concentration index of seismic source was constructed by the principal component analysis technique, to study the distribution characteristics of seismicities, and the variation rules of seismic energies and events numbers were also analyzed. The results show that: 1) Strong tremors mainly concentrated in the coal pillar regions formed by stop lines of multi coal faces in coal seam group; 2) Before strong tremor happened, the total energy of seismic events may be active firstly, then presents obvious decreasing trend, while the event numbers always keep in a relatively high level; 3)The larger the event number, and the smaller the concentration index of seismic source, the higher occurrence possibility of strong tremor. Based on the precursory rules of above parameters, the grading standard system of burst danger for No.10 Coal Mine of Pingdingshan Mining Area were built by accumulation method to evaluate the burst danger level. Application results show that the evaluation system can be successfully used to evaluate rock burst danger.

Coal bump inducing rule by dip angles of thrust fault

LI Shouguo, LYU Jinguo , JIANG Yaodong, JIANG Wenzhong

2014, 31(6): 869-875.

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In this paper, the main control parameters for coal bumps were generalized, and the mechanical mechanism of coal bump induced by fault dip. In addition, the simplified model of thrust fault was constructed, and the effect law of stress field, energy field and roof subsidence, etc., that inducing coal bump was simulated and studied in the conditions of different fault dip angles. The results show that when mining in the hanging wall, the peak abutment pressure, peak elastic energy and roof subsidence increase with the increase of fault dip in the condition of fault dip smaller than 45°, and while the fault dip is greater than 45°, the opposite results will be obtained. When mining in the footwall, the peak abutment pressure in front of the face decreases with the increase of fault angle. While the fault dip is 60° or 75°, the peak abutment pressure moves forward, which will raise the coal bump risk. The peak elastic energy and roof subsidence ahead of the face decrease with the increase of fault dip in the condition of fault dip smaller than 45°, and while the fault dip is greater than 45°, the opposite results will be obtained. Whether mining in hanging wall or footwall, the elastic energy and vertical stress in the fault area will increase with the increase of fault dip angle. In a word, fault dip angle has a greater influence to coal face when mining in the footwall.

Study on mechanisms of rock burst induced by a left coal pillar and prevention technology

YANG Weili, JIANG Fuxing, WEN Jinglin, LIU Yi, WEI Quande, YAO Shunli

2014, 31(6): 876-880.

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Taking left coal pillar with the width of 40 m in 5301 working face of a mine as the research background, the distribution characteristics of vertical stress has been obtained by means of theoretical calculation and numerical simulation. Based on this, the rock burst mechanisms are studied and the results are as follows: The impedance force of coal pillar is higher than impact force before disturbed by mining, so rock burst is less likely to occur. But in the stage of mining, the advance abutment pressure of working face is superposed by the original stress, consequently impact force is greater than resistance force and the risk of rock burst may increase greatly. According to this, the partition of dangerous areas and dangerous levels is obtained. The working face can pass by left coal pillar zones safely through implementing pressure relief and relief measures. The accuracy of the analysis result was proved by monitoring dynamic stress.

Evolutionary characteristics of mining stress near the hard-thick overburden normal faults

JIANG Jinquan, WU Quanlin, QU Hua

2014, 31(6): 881-887.

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Mining stress appears singularity under the influence of faults of hard thick overburden. By using three-dimensional numerical simulation method, the evolutionary characteristics of mining stress with the working face advancing to normal faults and layout along normal faults have been studied．Research shows that normal faults significantly weaken the mining stress transmission of hard thick overburden and hence, the prominent obstruction of mining stress. Roof fault zone is in a low stress state, while the stress is concentrated in the bottom fault zone. When hanging wall is advancing to normal faults, the overburden between working face and faults becomes“inverted wedge”shaped，and the mining stress is high，which becomes the key areas of disaster prevention. The lateral overburden of faults is“wedge”shaped, and mining stress is lower than that of the stress of the primary rock. When footwall is advancing to normal faults, the mining stress is low. Under the influence of fault cut, strong bearing area is formed on the lateral coal body of the end of working face. High stress concentration area appears in the front of faults coal pillar or working face when the hanging working face is distributed along the fault strike, and the stress concentration of surrounding rock in roadway is prominent, which makes it become a key harnessing zone.

Analytical solution of critical water inrush pressure of mining floor affected by fault

LU Haifeng, SHEN Dan, YAO Duoxi, HU Youbiao

2014, 31(6): 888-895.

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Considering the combined effect of mine pressure and water pressure of confined aquifer under the water-resisting floor, water pressure expression of floor water inrush has been derived. And then the most dangerous shear fracture plane and the critical pressure of water inrush have been obtained by trial calculation. Based on the proposed analytical solutions, the effect law of distance between cut-hole and fault zone, fault dip, advancing direction of face and lateral pressure coefficient on critical pressure of water inrush have been analyzed. The results show that the critical pressure of water inrush decreases gradually with the increase of distance between cut-hole and fault zone. With the decrease of fault dip, the critical pressure of water inrush decreases firstly, and then increases gradually. However, floor failure is not along the fault surface any more when the fault dip is less than a certain critical angle. When the working face advances from the lower wall of fault to the upper, the critical pressure of water inrush is higher than the opposite, namely the working face advances from the upper of fault to the lower. But no matter in what direction the working face advances, the critical pressure is reducing with the increase of distance from working face to the fault zone gradually. The critical pressure of water inrush increases first and then decreases with the increase of lateral pressure coefficient. The application demonstrates that the calculation results of the critical pressure accord with the actual situation.

Prediction of surface fissure in high relief areas induced by underground coal mining

HAN Kuifeng, KANG Jianrong, WANG Zhengshuai, WU Kan

2014, 31(6): 896-900.

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To objectively predict the location and width of surface fissures in high relief coal mining subsidence areas, according to the generation mechanism of surface fissures, a slope segment deformation has been predicted based on the vectors’ decomposition and synthesis principles using the equal gradient slope segment in the main section as research object and the subsidence and horizontal movement prediction values as main parameters. Taking the critical deformation value in rock and soil mechanics as the criterion of the beginning of surface fissures, this model is used to determine the location of the slope segment of the surface fissures and the width of surface fissures per unit length induced by mining. The feasibility and effectiveness of the method are tested by a case study of the 1208~1210 double face subsidence area of Jialequan Coal Mine in Shanxi Province, which shows a good fitness between the observed results and the predicted values and distribution pattern of slope segment deformation and location of surface fissures.

Asymmetric supporting research of gob-side entry driving in complex tectonic stress mining area

ZHOU Gang, WANG Pengju, ZOU Changlei, WANG Hao, WANG Chao

2014, 31(6): 901-906.

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When the working face is in the complex tectonic stress area，in order to determine the area ground stress condition，the method of stress relief is used to determine the region of high stress mainly constituted by tectonic stress. Based on the in-situ stress test results，the stress intensity and direction are used to load the initial ground stress in the numerical calculation model, which effectively reduces the error of the simulation results without considering the tectonic stress．Combined with evolution mechanism of deformation and stress along the empty roadway, asymmetric supporting parameters has been designed．Using special anchor fastening shoulder artifacts effectively can reduce the chances of fracture anchor cables and improve the quality of the anchor cable bearing quality．The results of the industrial test and the rock pressure indicates that the asymmetric support scheme supporting effect is good in the complex tectonic stress mining area and has a remarkable economic and social benefit．

Research on catastrophe instability mechanism of section coal pillars in block mining

CAO Shenggen, CAO Yang, JIANG Haijun

2014, 31(6): 907-913.

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A lot of residual coal pillars of irregular blocks coal resources were left after longwall mining. These coal resources are mainly exploited by the short-wall mechanized block mining technology, using continuous miner. Whether the section coal pillar is stable or not is the key to safety of irregular blocks coal mining. According to the geological production conditions of the 2-2 seam in Datong Coal Mine Group，catastrophe theory has been applied to analyze the catastrophe instability mechanism of section coal pillars. The essential condition for catastrophe instability of section coal pillars has been derived: when plastic zone width of coal pillar is more than 86%. The numerical simulation results will conform to the observations. The dynamic evolution process of failure in section coal pillars and stope has been simulated . Taking the section coal pillar in Datong Coal Mine as an example, the theoretical calculation results for catastrophe instability is that plastic zone width of coal pillar should be no more than 12.9 m, which is consistent with drilling peep experimental results.

Study on technology of gob-side entry retaining in thin seam surrounded by soft roof and floor

JU Feng, SUN Qiang, HUANG Peng, HE Qi

2014, 31(6): 914-919.

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In response to the special mining condition of the weak roof and floor in the working face of the thin seam, the difficulties and key technology of gob-side entry retaining in thin seam have been proposed. By using the theory analysis and numerical simulation, the failure characteristics and deformation law of surrounding rocks of gob-side entry retaining in thin seam have been analyzed. The structure model of gob-side entry retaining has been set up in thin seam coal mining. At the same time, Through the numerical simulation analysis and expounding the basic principle gob-side entry retaining in thin seam, it has been put forward that the optimization scheme of gob-side entry retaining was to “built waste stone wall in the goaf first, then timely build cemented filling body and long steel beams and monomer pillar support in roadway at last”. Accordingly, the gob-side entry retaining speed was improved obviously. The result of the actual measurement in Zhaoguan coal mine shows that the roof subsidence is only 197 mm; the floor heave is 83 mm；the convergence of two ribs is 327 mm, consequently, the mining efficiency reached up to 18 m/d.

Stability control mechanism of mudstone wedge structure roof in crossing coal seam roadway

JI Ming, GUO Hongjun, LIU Haiquan, ZHANG Yidong, CHENG Liang, WANG Gang

2014, 31(6): 920-925.

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The main controlling factors for stability of roadway roof in wedge model have been proposed in response to the structural characteristics of crossing coal seam roadway; meanwhile, mechanical model of roof cantilever beam for wedge segment has been established. Through static strength analysis and deflection calculation of this roof section, the formula for solving the minimum support resistance was obtained. By imputing engineering and geological conditions of Gucheng Mine of Lu’an Group into the resistance formula, it is concluded that the minimum support resistance in the model is 0.12 MPa. Corresponding anchor cable supporting scheme has been proposed considering safety factor as 1.4. Field industrial experiments and mine pressure monitoring results show that the supporting design scheme based on calculating support resistance can control roof deformation and maintain roadway stability.

Applied research on a hole-shaped support model for pressure relief in heavily stressed soft rock roadway

LI Yanbin, LI Ligong, LI Jiankun, MA Dong

2014, 31(6): 926-931.

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Through the analysis of failure mechanism of high stress soft rock tunnel excavation, the hole-shaped support model for pressure relief (surrounding rock-material of pressure relief- steel bracket) has been built, the supporting principle of pore pressure relief has been analyzed, and the support pattern for pressure relief in heavily stressed and soft rock roadway has been put forward . By making use of the software of RFPA, the evolutionary process of surrounding rock failure before and after the supporting in the rectangle roadway has been imitated. The result indicates that, when there is no support in the roadway, there is overall damage with fractures in both sides mainly ; when using steel shed support, the damage in the sides and bottoms will induce a wide range of roof abscission layer, rock falling and collapse of the steel shed; when adopting the hole-shaped support for pressure relief, surrounding rock stress can release through the compression of the pressure relief materials . So there is no obvious deformation in roadway surface, and the support is lossless, which indicates that this kind of support can reduce the surrounding rock stress, and enhance safety of roadway. It provides a new method of supporting for the heavily stressed soft rock roadway.

Research on influencing factors of excavation disturbance

HAN Jiaguang, YANG Sen, ZHANG Nong

2014, 31(6): 932-937.

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In deep mining, the excavation disturbance is large. Such factors of excavation disturbance as strength of surrounding rock of roadway, horizontal distance, vertical distance have been discussed in this paper, adopting monitoring and numerical simulation results. The result indicates that the increase of the rock internal friction angle and cohesion can weaken disturbance effect of the roadway, while tensile strength is almost not affected to the excavation disturbance; in the roadways when the distance exceed 5 times of theirs diameter condition, compared to vertical distance, the increase of the horizontal distance can significantly weaken the excavation disturbance of soft rock roadways; In vertical direction, The distribution of the excavation disturbance effect is different accordingly. When early roadway is located above the later roadway, increase in the horizontal distance can weaken the excavation disturbance effectively. On this basis, the determination method is proposed for optimal spatial relationship of roadways, which has certain significance on engineering practice.

Study on strengthening control mechanism with support and cable for roadway in“three-soft”coal seam during deep mining

JING Shengguo, WANG Qizhou, CHEN Jie

2014, 31(6): 938-944.

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In response to the problem of the unstability of the “three-soft” coal roadways caused by serious deformation despite the repeated reinforcement of it in Zhengzhou mining area, the characteristics of “three-soft” coal seam occurrence, stress environment and the causes of the unstability have been analyzed. On this basis, the strengthening control mechanism with support and cable has been proposed. In this supporting system, on one hand, the bearing capacity and stability of support are improved by the supporting resistance from the cable. On the other hand, the shear deformation of creep coal seam is effective controlled by higher supporting resistance of support. In addition, the bearing capability of coal seam is also developed, which makes the collaborative bearing of the support, cable and surrounding rock possible, and according to the mechanical calculation, the primary and secondary parts to be strengthened and the reasonable determination of the cable prestress are indicated. After the application of support and cable strengthening technology in 31041 haulage roadway of Chaohua mine, convergence of two sides in the roadway are controlled less than 300 mm during the impact of working face mining, and convergence between roof and floor are controlled less than 400 mm, which can control effectively the deformation of deep “three-soft” coal roadway.

Failure mechanism and constant resistance large deformation control measures of deep soft rock in Qingshui Coal Mine

GUO Zhibiao, WANG Jiong, ZHANG Yuelin, PAN Qixin, PENG Linjun

2014, 31(6): 945-949.

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With mining depth increasing, deep roadways emerge serious floor heave and roof convergence by complex stress field. Softrock roadways deformation of Qingshui Coal Mine is the most severe in Shenbei Mining Area. By researching on lithology and stress field, the paper draws reasons and failure mechanism of the typical roadway which use the traditional supporting methods. The deep soft rock supporting technology which is constant resistance large deformation bolt coupling technology was put forward. The technology control roadway deformation by constant resistance and energy absorption. The support method works well on site which provides a new way of control deep soft rock roadway large deformation.

The research on stability and supporting technology of rock in gob-side entry in thick seam with parting

BUI Manhtung, LU Yan, GUO Weibin, WANG Feilong, ZHAO Zhanquan

2014, 31(6): 950-956.

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In order to ensure the stability of the thick coal seam with parting in gob-side entry in longwall top coal caving, and to increase the safety and effectiveness in mining, the numerical model has been employed to analyse the stress distribution of gob-side entry, plastic strain area distribution and movement of rocks under the different conditions of dirt bands. Result of research shows that the existence of band layers has destroyed the continuity of the surrounding rock, which causes the roads becoming unstable; while in the thick layers, surrounding rocks are relatively stable. When hardness of the layers and that of coal seam are the same, and these bands lay on the top of road, the rocks surrounding the roads are also relatively stable. And on this basis, an appropriate technique of surrounding rock control is proposed, in which a combination of rock bolting-wire mesh and steel rectangular shape support are chosen and the suitable parameters are identified. Engineering practice has proved the rationality of the proposed technique, ensuring the safety and effectiveness in long wall extraction.

Numerical analysis of split Hopkinson pressure bar test with passive confined pressure for coal

LI Chengwu, WANG Jingui, XIE Beijing, HU Bo, WANG Chuan

2014, 31(6): 957-962.

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Coal or rock dynamical disaster poses grave threats to the coal mine safety and high efficiency production. So, research on dynamic properties of coal under the complex stress state is significant. By using the software LS-DYNA, the Φ50 mm split Hopkinson pressure bar(SHPB) tests for coal have been numerically simulated by using the Holmquist-Johnson-Cook (HJC) constitutive model. Based on the finding that the numerical simulation results were similar to the experimental results, passive confined pressure SHPB of coal has been simulated. It is demonstrated that elasticity modulus of sleeve, thickness of sleeve, friction coefficient of contact surface, interval between sleeve and specimen significantly have significant impact on dynamic properties of coal. According to the research results, a new layer to the coal pillar can be added. It can improve shock resistance of coal pillar by designing optimal parameters of additional layer: elasticity modulus, thickness, friction coefficient of contact surface interval.

Numerical simulation research of splitting failure of roadway under high stress

NIE Taoyi, PU Hai

2014, 31(6): 963-968.

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In response to the splitting failure which is a special engineering geological phenomenon under high in-situ stresses, splitting model based on curve expanding path involving closure effect has been applied. This model is programmed by FISH functions and embedded into FLAC software. Numerical simulation of splitting failure of roadway has been conducted by using FLAC software. The diagrams of splitting failure evolution with respect to the depth and area under different geological conditions have been obtained. The impacts of depth, lateral pressure coefficient, bulk modulus and shear modulus on splitting failure have also been analyzed. The research can offer reliable references to the evaluation of stability of underground structures and the design of support under high in-situ stresses.

Research on the deformation features and strengthening technology in auxiliary shaft lining under the complex stress

CHEN Xiaoxiang, XU Yichang, FAN Zengzhe, LIU Jun

2014, 31(6): 969-975.

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Based on the engineering background of the auxiliary shaft of Shihao mine of Henan Da- you Energy Co Ltd, by field investigation and research, deformation and failure features and damage reasons of deformation zone in mine shaft have been analyzed, by using FLAC3D numerical simulation software, the stress distribution and the plastic zone size inside of rock pillars on both sides of shaft and the displacement variation rule of shaft lining on side of two rock pillars have been studied under the different pre-tightening force of anchor rope. The research results show that under the action of the strand anchor, as the pre-tightening force of anchor cable is increasing, vertical stress within the rock pillars will increase, while the plastic zone size and the shaft displacement after reinforcement will decrease. Self-supporting ability and stability of shaft lining surrounding rock are effectively improved by the action of high pre-tightening force of the strand anchor. Full grouted anchor cable can effectively reduce the post loss of anchor pretightening force, and then the wall rock stability is enhanced. This numerical simulation results were better validated by shaft lining surrounding rock deformation measured results. The research results provide important references for treatment rupture of the shaft lining under similar conditions.

An analysis of the development of artificial freezing wall in seepage strata with fractured structure and its influencing factors

FENG Meimei, YANG Weihao, GAO Juan

2014, 31(6): 976-981.

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A numerical calculation model of fractured rock mass frozen with double hollows under the function of seepage has been set up，and the influence of the fracture aperture, the fracture inclination and the influence of the seepage velocity on the development of artificial freezing wall in seepage strata with fractured structure have been analyzed using the COMSOL Multiphysics finite-element analysis software．The results show that: 1) The larger the fracture aperture is, the longer the freezing wall closure time becomes and the larger the increase; both in the main section and the interface of frozen wall, the upstream speed of development in the frozen wall is significantly less than that of the downstream speed of development; 2) The freezing zone and isotherms of fracture angle of 0 °are symmetric about the y-axis, and the distribution of freezing temperature field for 10 °, 20 ° and 30 ° is no longer symmetrical about the y-axis, and the isotherms outside the frozen wall show “heart-shaped” distribution along the fracture inclination, ie the thickness of the frozen wall is the thickest in the downstream direction while the thinnest in the upstream direction along the fracture inclination. The influence of water seepage on the freezing temperature field distribution decreases gradually as the fracture angle increases. Under the given conditions in this paper, the seepage velocity is the main factor which influences on the frozen wall development, followed by fracture aperture and the fracture inclination.

In-situ measurement and study of freezing pressure of shaft in western cretaceous water-rich bedrock

YANG Gengshe, QU Yonglong, XI Jiami, CHEN Xinnian, LI Borong

2014, 31(6): 982-986.

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To solve the technological problems of mine construction by freezing of cretaceous water-rich bedrock in western region, by using the vibrating wire sensors to measure and analyze bedrock’s freezing pressure and interface temperature of air shaft in a Gansu mine, the change rules, causes and inhomogeneity on freezing pressure of cretaceous bedrock during shaft sinking have been studied. The results show that the freezing pressure contains four stages: rapid growth, dramatic increase, slow rise and gradual steadiness. It has obvious inhomogeneity at the same level and is caused by many factors. Its final stable value reaches 1.708-2.047 MPa, less than the maximal theoretical and empirical value of western bedrock at the same depth, and the upper limit of freezing pressure pω in eastern shock layer. The interface temperature of frozen wall and shaft lining rises sharply and then reduces quickly, with the maximum of each point reaching 42.4-59.4 ℃ within 30 h and the difference values 51.86-71.3 ℃after the temperature decrease, which is extremely unfavorable to the security and stability of the two wall. Following the eastern design methods and experiences tends to cause the wall to become much thicker and produce ringed temperature cracks because of temperature stress, so there still exists an optimized space on structural design and construction of the western mine shaft.

Seepage characteristics of coal mass containing gas considering moisture effect in loading-unloading confining pressure test

WEI Jianping, QIN Hengjie, WANG Dengke

2014, 31(6): 987-994.

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In this paper, taking coal samples in Fangshan Coal Mine, Henan as research object, the self developed “Loaded coal containing gas triaxial loading seepage experiment device” was used to conduct the loading-unloading confining pressure triaxial seepage experiments under the condition of different moisture contents, and to analyze the influence law of moisture and loading-unloading confining pressure on the seepage characteristics of coal mass containing gas. The results show that: 1) Confining pressure and moisture content of coal sample have strongly control ability to the permeability of briquette, while the permeability is correlated negatively with the moisture and confining pressure. 2) In the two loading processes, the permeability variation in the second loading process is gentler, and the relation between permeability and confining pressure curve path is different. 3) During the two unloading processes, the permeabilities increase to some extent, but can’t recover to the initial value. The higher the moisture content is, the smaller the recovery degree is. 4) During the same loading or unloading process, the permeability of coal sample with higher moisture content changes more gently. After the first loading, during the first unloading and the second loading-unloading processes, the control ability of moisture content in the permeability of coal sample is much higher than that of the confining pressure. Experimental conclusions can provide reference to the choice of the main control factors to gas permeability in coal seam under the repeated loading-unloading conditions.

Research and application of gas drainage with hydraulic reinforcement of water stimulation fracture in borehole through strata

XIN Xinping, GAO Jianliang, MA Geng

2014, 31(6): 995-1000.

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In order to solve the problem of how to efficiently extract the coal seam with low permeability, the hydraulic measures of stimulation fracturing through strata borehole has been adopted to improve the permeability of coal seam. The gas migration process through borehole in the hydraulic enhanced technology has been analyzed. The effect of hydraulic reinforcement measures on hard and soft coal seam permeability has been studied, and the effect and mechanism of hydraulic reinforcement on the permeability of coal mass has been analyzed, and finally, the field experiment of hydraulic fracturing has been carried out. Research results show that the hydraulic measures of stimulation fracturing through strata borehole technology can improve the permeability of coal seam and contribute to form gas flow channel. There is “virtual reservoir” in roof and floor in hard coal seam and the gas shows seepage mainly, while, in soft coal seam gas shows diffusion migration mainly. The hydraulic measures of stimulation fracturing can increase the permeability of coal seam through formation of cave and cracks in it, release ground stress and reduce gas flow attenuation coefficient, and consequently improve efficiency of the gas drainage. By using hydraulic fracturing measures in coal seam the pure quantity of gas drainage increases from 14.8 m3 to 31.9-42.8 m3; the gas concentration increases from 14.1% to 73.9%-77.5%. This technology has very wide popularization and application prospects.

Exceptional response characters of radio wave perspective in coal thinning zone

JIAO Xianfeng, JIANG Zhihai, LIU Shucai

2014, 31(6): 1001-1004.

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Accurate delineation of the coal thinning zone caused by igneous rock intrusion, is beneficial to the high yield and high efficiency for the mechanized coal mining face. The igneous rock intrusion has been detected by the radio wave perspective on one coal face. The initial amplitude and the absorption coefficient have been got by comprehensive curves analysis to reconstruct the absolute attenuation tomography. The absorption coefficient β can denote the exceptional response features. The contrast between tomography results and the actual coal seam thickness shows that: 1) The anomaly of coal thinning zone caused by igneous intrusion is obvious in absolute attenuation tomography map. Its distribution range conforms very closely to the perspective anomaly area, and radio wave perspective to detect the coal thinning zone is feasible. 2) The absorption coefficient β is corresponding to the igneous rock erosion thickness. According to the difference between abnormal absorption coefficient and normal absorption coefficient, the coal thickness can be determined.

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