Article
Soft Roof Failure Mechanism and Supporting Method for Gob-Side Entry Retaining Hongyun Yang 1,2 , Shugang Cao 1,2, *, Yong Li 1,2 , Chuanmeng Sun 1,2,3 and Ping Guo 4 Received: 12 May 2015 ; Accepted: 20 October 2015 ; Published: 28 October 2015 Academic Editor: Saiied Aminossadati 1 2 3 4
*
State Key Laboratory of Coal Mine Disaster Dynamics and Control, Chongqing University, Chongqing 400044, China;
[email protected] (H.Y.);
[email protected] (Y.L.);
[email protected] (C.S.) College of Resources and Environmental Science, Chongqing University, Chongqing 400044, China School of Computer Science and Control Engineering, North University of China, Taiyuan 030051, China National Key Laboratory of Gas Disaster Detecting, Preventing and Emergency Controlling, China Coal Technology Engineering Group Chongqing Research Institute, Chongqing 400037, China;
[email protected] Correspondence:
[email protected]; Tel./Fax: +86-23-6511-1706
Abstract: To study the soft roof failure mechanism and the supporting method for a gateway in a gently inclined coal seam with a dip angle of 16˝ kept for gob-side entry retaining, and through the methodology of field investigation and numerical and analytical modeling, this paper analyzed the stress evolution law of roof strata at the working face end and determined that the sharp horizontal stress unloading phenomenon along the coal wall side did not appear after the working face advanced. Conversely, the horizontal stress along the gob side instantly decreased and the tensile stress produced, and the vertical stress in the central part of the roof had a higher reduction magnitude as well. An in-depth study indicates that the soft roof of the working face end subsided and seriously separated due to the effect of the front abutment pressure and the roof hanging length above the gob line, as well as certain other factors, including the rapid unloading of the lateral stress, tension and shear on the lower roof rock layer and dynamic disturbance. Those influencing factors also led to rapid crack propagation on a large scale and serious fracturing in the soft roof of the working face end. However, in the gob stress stabilized zone, the soft roof in the gob-side entry retaining has a shearing failure along the filling wall inside affected by the overburden pressure, rock bulking pressure, and roof gravity. To maintain the roof integrity, decrease the roof deformation, and enable the control of the working face end soft roof and the stabilization of the gob-side entry retaining roof, this study suggests that the preferred bolt installation angle for the soft roof situation is 70˝ based on the rock bolt extrusion strengthening theory. Keywords: gob-side extrusion strengthening
entry
retaining;
unloading
loose;
bolt
installation
angle;
1. Introduction The retaining of the gob-side entry is to maintain the head entry of current mining panel behind the working face to be reused for the next panel as the tail entry. This technology can effectively increase the coal recovery rate, reduce the roadway development rate, and mitigate the outburst risk, as no pillar is needed for the retained entry, but rather an artificial filling wall is constructed on the gob side with a special support for entry stability [1]. Since the 1950s, pillarless roadways have been constructed worldwide, and extensive studies have been carried out on support resistance [2,3]. Previous studies showed that the gateway roof rock mass failure mechanism and supporting methods have a significant impact on gob-side entry Minerals 2015, 5, 707–722; doi:10.3390/min5040519
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retaining, determining the procedure’s success. Many researchers categorized five failure types of a gateway roof rock mass, including the compressive stress failure type, tensile stress failure type, shear stress failure type, squeezing and fluidity failure type, and geological structure failure type [4–6]. Based on the above failure mechanism of gateway roof strata, many active support models were suggested, and a combined supporting technology using a bolt, mesh, and cable was found to be effective for ground control. It was found that a combined support is unable to control the main roof in roadway retaining according to the theory of “given deformation”, but it can stabilize the immediate roof with the main roof to maintain the retained gob-side entry very well [7,8]. Rock bolts have become a popular technique for reinforcing rock masses all over the world. Rock bolts are installed to reinforce a fractured rock mass by resisting dilation or shear movement along the fractures. Nemcik et al. [9] determined that the non-linear bond-slip relationship used in the FLAC model for bolting accurately matches the experimental data reported by Ma et al. [10]. Yang [11] divided the anchored force evolution process into three stages and found that a surface structure-like bearing plate produces a maximum anchorage force acting directly on the surface of rock that can provide the greatest degree of support. This not only improved the rock mass stress state but also increased the thickness of the reinforcement structure formed by the anchoring force. Zheng and Zhang [12] determined the shear stress and pressure stress distribution equation for the anchored segment and upon studying the stress distribution rule of anchored segment, found that the effect of the pre-tensioned bolt in the soft rock was superior to that in hard rock. Cao [13] showed that the integrity of the supporting system that prevents local failure of surrounding rock from progressing into overall failure is important in rock bolting, so that reinforcing measures should be taken when necessary. Fan [14] designed a roadway heterogeneity controlling technology on a siltstone roof. Hua [15] made use of bolt support and anchor cable reinforcement support technologies inside and beside a retained roadway, respectively, and maintained the retained gob-side entry very well. Yan [16] used pre-tensioned bolts on a roadway roof with medium and fine-grained sandstones by an extrusion lockset to limit its vertical deformation for entry integrity and also analyzed the mechanical mechanism, technical principle and technical characteristics of bolt and cable coupling supports at a mine site. Chen and Bai [17] used bolts with high pre-tensioning, high strength, and a large elongation rate and cables as the basic supporting element, a single hydraulic prop with a metal hinge top beam as a reinforcement support inside the roadway, and a high-water rapid-solidifying material to support the side of the roadway to effectively control the roof rock deformation with medium and fine-grained sandstones in gob-side entry retaining. In the past experimental studies and supporting theory [4,14–17], the bolt support is mainly aimed at the control of the medium-hard roof rock mass, which are generally medium and fine-grained sandstones in the horizontal and flat seams. The supporting theory also provides supporting references for the similar conditions of the roadway in the initial excavation stage and retaining stage. However, the working face end roof and retained roadway roof seriously deformed in the soft roof rock mass and gentle inclined coal seam with a soft roof rock after referencing the existed supporting theory. Therefore, to further develop our understanding of the soft roof failure mechanism and provide a supporting method, the following works were conducted in this article. First, the roof failure phenomenon of a gently inclined coal seam working face end and retained roadway were investigated at a coal mine in the southwest of China. Second, the roof failure mechanism was analyzed through the field investigation and stress evolution law obtained by three-dimensional distinct element code (3DEC). Finally, a roof support was designed based on results of this investigation to maintain roadway stability. 2. Description of Field Observation 2.1. Survey of Study Site The target coal mine is located in the southwest of China. The mining coal seam is #C19, with an average dip angle of 16˝ and a mining thickness of 1.1 m. The mining method is the fully mechanized 708
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longwall, and gob side entry retaining is used, as shown in Figure 1a. The investigated head entry is buried at 315 m, with a cross-section of 3.8 m in width and 2.1 m in height of the short rib, where Minerals 2015, 5, page–page resin-anchored rock bolts (Φ20 ˆ 2200 mm), cables (Φ12.54 ˆ 6300 mm), and reinforcement mesh were and three of single hydraulic supports with tophinged hingedbeams beamswere were used used as used,used, and three rowsrows of single hydraulic supports with top as advanced advanced strengthening support, as shown in Figure 1b. The roof rock bolts are non-fully anchored strengthening support, as shown in Figure 1b. The roof rock bolts are non-fully anchored (anchor (anchor length 1.4 m) m) with with aapre-tension pre-tensionofof90 90kN, kN,inter-row inter-rowspacing spacingofof 800 ˆ 800 angles of length is is 1.4 800 × 800 mmmm andand boltbolt angles of 90°, ˝ , 80˝ , 70˝ , and 60˝ for different bolts, and the extra cables have a pre-tension of 200 kN and an 90 80°, 70°, and 60° for different bolts, and the extra cables have a pre-tension of 200 kN and an interinter-row spacing of 3200 ˆ mm. 1200 mm. row spacing of 3200 × 1200
Figure 1. (a) Schematic of gob-side entry entry retaining; retaining; (b) (b) Schematic Schematic of ofadvanced advancedstrengthening strengtheningsupport. support.
According According to to aa geological geological survey, survey, the the gateway gateway roof roof strata strata typically typically consist consist of of weak weak rocks. rocks. The The immediate immediate roof roof is is approximately approximately 1.0 1.0 m m in in thickness thickness and and is is made made up up of of sandy sandy mudstone mudstone and and mudstone. The main roof consists of more competent siltstones and silty mudstones containing mudstone. The main roof consists silty mudstones containing siderite Inother otherwords, words,the the roof strata have strength stability and exhibit siderite nodules. In roof strata have lowlow strength andand poorpoor stability and exhibit clear clear stratification, resulting an immediate roof caving the working advancing as stratification, thus thus resulting in aninimmediate roof caving with with the working face face advancing as well well as part of the main roof. Also, the floor is made up of sandy mudstone. It must be stressed as part of the main roof. Also, the floor is made up of sandy mudstone. It must be stressed that, at the that, at the working face end, thesoft gateway softoften roof affected was oftenbyaffected byloading dynamicforloading for the working face end, the gateway roof was dynamic the following following reasons: caving of the overburden mass the gob and themovement, gangue movement, reasons: (1) caving(1) of the overburden rock massrock in the gobinroof, and roof, the gangue collision; collision; (2) periodic weighting the main blasting operationsofofadjacent adjacent mining mining and (2) periodic weighting of the of main roof; roof; and and (3) (3) blasting operations and excavating excavating faces. faces. 2.2. 2.2. The The Survey Survey Results Results of of Soft Soft Roof Roof Failure Failure Characteristics Characteristics The The soft soft roof roof failure failure characteristics characteristics of of the the front front abutment abutment pressure pressure zone, zone, the the stress stress decreasing decreasing zone, and the stress stability zone are shown in Figure 2. According to a filed investigation zone, and the stress stability zone are shown in Figure 2. According to a filed investigation of of the the head entry, there was a serious roof deformation due to the impact of front abutment pressure anda head entry, there was a serious roof deformation due to the impact of front abutment pressure and afew fewcracks crackswith withless lessopening opening over a small area roof surface, shown in Figure With over a small area of of thethe roof surface, as as shown in Figure 2a.2a. With the the working face advancing, the roof overhanging length increased, and cracks formed and joined working face advancing, the roof overhanging length increased, and cracks formed and joined together together in in the the roof roof rock rock mass, mass, as as shown shown in in Figure Figure 2b, 2b, resulting resulting in in aa decrease decrease of of the the roof roof strength, strength, loss loss of self-bearing capacity, separation from the main roof, and roof deformation and becoming unload of self-bearing capacity, separation from the main roof, and roof deformation and becoming unload rock rock body, body,especially especiallyfor forthe therock rockmass massindicated indicatedby bythe thered redline. line. In In addition, addition, two two bolts bolts and and one one cable cable (B1, B2 and C1) on the gob side lost their support ability. After building the filling wall and removing (B1, B2 and C1) on the gob side lost their support ability. After building the filling wall and removing the the supports supports in in the the stress stressstability stabilityzone, zone, the the roof roof fractured fracturedalong alongthe theinside insideof ofthe thefilling fillingwall, wall,indicated indicated as the “fracture line” in Figure 2c. as the “fracture line” in Figure 2c. Field Field observations observations found found that that there there were were aa large largenumber number of of cracks cracks in in the the vertical vertical direction direction of of roof roof near the end of the working face, so the roof caving occurred generally within a height range of near the end of the working face, so the roof caving occurred generally within a height range of 1.5 1.5 to to 22 m, m, as as shown shown in in Figure Figure 3. 3.
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Figure 2. Schematic of soft roof rockmass massfailure; failure; (a) (a) In pressure zone; (b) In Figure 2. Schematic of soft roof rock In front frontabutment abutment pressure zone; (b)working In working Figure 2. Schematic soft roofzone rockofmass failure; (a) In front abutment pressure zone; (b) In working face end; (c) In stressofstability retaining gateway. face end; (c) In stress stability zone of retaining gateway. face end; (c) In stress stability zone of retaining gateway.
Figure 3. Caving roof near the working face end. Figure 3. Caving roof near the working face end.
Figure 3. Caving roof near the working face end. 3. Stress Evolution Law in Roof 3. Stress Evolution Law in Roof analyze the stress evolution law in the roof, the numerical modeling was adopted. 3. Stress To Evolution Law in Roof To analyze the stress evolution law in the roof, the numerical modeling was adopted.
To the stress evolution law in the roof, the numerical modeling was adopted. 3.1.analyze Numerical Simulation Model 3.1. Numerical Simulation Model
Because of the coal seam condition and as the rock mass was discontinuous [18,19], 3DEC of 3.1. Numerical Simulation Model
Because of the coalMN, seam condition thetorock mass discontinuous 3DECThe of ITASCA (Minneapolis, USA) [20,21]and was as used study thewas stress evolution law[18,19], in the roof.
ITASCA (Minneapolis, USA) [20,21] was used study the m, stress law in [18,19], the roof. The of Because of model the coal condition and as the rock was discontinuous 3DEC hexahedral has seam aMN, length, width, and height ofto230 m,mass 100 andevolution 200 m, respectively, includes hexahedral model has a length, width, and height of 230 m, 100 m, and 200 m, respectively, includes coal seams and rock strata, with a total of 14 layers, as shown in Figure 4, in accordance with the The ITASCA (Minneapolis, MN, USA) [20,21] was used to study the stress evolution law in the roof. coal seamsconditions and rock strata, with a total of 14 layers, as shown in Figure 4, in accordance with the geological of the investigated mine. The Mohr-Coulomb yield criterion for the materials hexahedral model has a length, width, and height of 230 m, 100 m, and 200 m, respectively, includes geological conditions of the investigated mine.used. The The Mohr-Coulomb yield criterion for the materials and theand Coulomb model foracontact were physical of allwith the the coal seams rock slip strata, with total of 14 layers, asmechanical shown inand Figure 4, inproperties accordance and the Coulomb slip model for contact were used. The mechanical and physical properties of all the layers and the contacts between every two layers are described in [13,22], respectively. geological conditions of the investigated mine. The described Mohr-Coulomb yield criterion for the materials layersAccording and the contacts between every two layers arecoefficient [13,22], respectively. to the reference [23], the minimum ofinlateral pressure is very close to 0 and and the Coulomb slip model for contact were used. The mechanical and physical properties all the Accordingcan to the reference [23], minimum coefficient of lateral is very to 0 of and the maximum be up to 6 due to the tectonic movement. For this case, pressure we consider the close coefficient of layersthe and the contacts between every two layers are described in [13,22], respectively. maximum be up due to tectonic movement. Fordepth this case, we consider the coefficient of lateral pressurecan is 0.5, as to the6 gateway is placed at shallow and close to the anticline axis and According to the reference [23], the minimum coefficient of lateral pressure is very close to lateral pressure is 0.5, as the gateway is placed at shallow depth and close to the anticline axis and seam outcrop, which resulting in a much low horizontal stress. So, the state of the in situ stresses is0 and seam resulting in muchwith low stress. So, the state of the direction in situ the maximum can which be upand to 6σzdue toaMPa, tectonic movement. Forthethis case, we consider thestresses coefficient σx = σoutcrop, y = 4.25 MPa = 8.5 σhorizontal y parallel to longwall advance and σisx of σ x = σ y = 4.25 MPa and σ z = 8.5 MPa, with σ y parallel to the longwall advance direction and σx and lateral pressure is 0.5, as theingateway is A placed atpressure shallowofdepth andis close to on thethe anticline axis perpendicular, as shown Figure 4b. vertical 8.5 MPa applied top surface, perpendicular, as shown in Figure 4b. A vertical pressure of 8.5 MPa is applied on the top surface, the velocity the bottominsurface was restricted in all stress. three directions, vx = vof y = the vz = in 0 m/s. seamand outcrop, whichofresulting a much low horizontal So, the state situ The stresses of the bottom surface was restricted innormal all threedirection directions, x m/s. = vy = vz = 0 m/s. The velocity of theMPa other four were restricted in the v n =v 0 is σx and = σythe = velocity 4.25 and σsurfaces = 8.5 MPa, with σ parallel to the longwall advance direction and σx z y velocity ofshape the other fourof surfaces were restricted in the normal direction vn = 0 an m/s. The and size the head entry are introduced in Section 2.1. After initial equilibrium perpendicular, as shown in Figure 4b. A vertical pressure of 8.5 MPa is applied on the top surface, The shape size of the head entry are introduced in once Section After initial equilibrium calculation, rockand bolts were installed, as shown in Figure 1b, the2.1. head wasan entry excavated. The and the velocityrock of the bottom surface was restricted in all three directions, vxentry = vyexcavated. = vz = 0 m/s. The calculation, bolts were installed, as shown in Figure 1b, once the head was The rock bolts were represented as built-in “cable” elements. For resin-grouted rock bolts, the stiffness velocity of the were otherrepresented four surfaces were restricted in the normal direction v = 0 bolts, m/s. rock as built-in elements. (Kbondbolts ) and the cohesive strength (Sbond) of“cable” the grout are the For two resin-grouted key propertiesnrock that governthe thestiffness anchor The shape and size of the head entry are introduced in Section 2.1. After an initial equilibrium (K bond ) and the cohesive strength (S bond ) of the grout are the two key properties that govern the 9 5 characteristics [22]; Kbond = 3.06 × 10 N/m/m and Sbond = 2.3 × 10 kN/m were adopted in thisanchor study. 9 N/m/m calculation, rockabolts shown Figure 1b,×modulus once theofhead was entry excavated. characteristics [22];were Kbond =installed, 3.06 × 10 = 2.3 105 kN/m were adopted this study. −4in In addition, cross-sectional area of as 3.142 × 10and m2S , bond an elastic 200 GPa, and aintensile yield The −4 m2, an elastic modulus of 200 GPa, and a tensile yield In addition, a cross-sectional area of 3.142 × 10 rock bolts were represented as built-in “cable” elements. For resin-grouted rock bolts, the stiffness (Kbond ) and the cohesive strength (Sbond ) of the 4grout are the two key properties that govern the 4 anchor characteristics [22]; Kbond = 3.06 ˆ 109 N/m/m and Sbond = 2.3 ˆ 105 kN/m were adopted in this study. In addition, a cross-sectional area of 3.142 ˆ 10´4 m2 , an elastic modulus of 200 GPa, and
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a tensile yield strength of 165 kN were assigned to the “cable” element. A stepwise excavation in the Minerals 2015,of 5, 165 page–page strength kN were assigned to the “cable” element. A stepwise excavation in the y-direction y-direction was adopted to simulate theface working face by in deleting blocks ineach, five as steps of was adopted to simulate the working advancing byadvancing deleting blocks five steps of 10 m ofshown 165 kNin assigned 10 mstrength each, asin Figure 4c. to the “cable” element. A stepwise excavation in the y-direction shown Figure 4c.were was adopted to simulate the working face advancing by deleting blocks in five steps of 10 m each, as shown in Figure 4c.
Figure 4. (a) Numerical model; (b) Coal seam mining model; (c) Working face advancing;
Figure(d) 4.Immediate (a) Numerical model; (b) Coal seam mining model; (c) Working face advancing; (d) roof grid, bolt and monitoring points; (e) Gateway cross section of model. Immediate roof grid, bolt and monitoring points; (e) Gateway cross section of model. Figure 4. (a) Numerical model; (b) Coal seam mining model; (c) Working face advancing; (d) Immediate roof grid, bolt and monitoring points; (e) Gateway cross section of model.
3.2. Numerical Results
3.2. Numerical Results To understand the stress evolution law, the horizontal stresses in the x-direction of points A1 and and the vertical of evolution point A2 in the y = 45 m, as stresses shown inin Figure 4d, were monitored 3.2. Numerical Results ToA3understand thestress stress law,roof theathorizontal the x-direction of points A1 during the working face advancing, and the results are shown in Figure 5. Especially, it can be found and A3 and the vertical stress of point A in the roof at y = 45 m, as shown in Figure To understand the stress evolution law, and were 2 the horizontal stresses in the x-direction of points A1 4d, that from A to B Zone, the horizontal stress on the coal side (A3) increases slightly in magnitude monitored thestress working faceA2advancing, are in shown 5. Especially, A3 and during the vertical of point in the roof atand y = the 45 m,results as shown Figurein 4d,Figure were monitored and then reduces to approximately 5.9 MPa, while the horizontal stress on the gob side (A1) drops the working faceAadvancing, and the results arestress shownon in the Figure 5. side Especially, it can be found it canduring be found that from to B Zone, the horizontal coal (A ) increases slightly in 3 instantaneously and appears tensile, and the vertical stress at the gateway central (A2) reduces that from A to B Zone, the horizontal stress on the coal side (A 3) increases slightly in magnitude magnitude and then reduces to approximately 5.9 MPa, while the horizontal stress on the gob side significantly to 4.5 MPa as soon as the working face advances beyond the monitor points. The stresses and then reduces to approximately 5.9 MPa, while the horizontal stress on the gob side (A 1) drops (A1 ) drops and over appears tensile, and from the vertical stress ofatthe thecaving gateway in theinstantaneously gateway roof changed the whole process the beginning to thecentral roof (A2 ) instantaneously andaappears tensile, the vertical stressinfluence at the gateway centralthe (A2gob-side ) reduces stabilization, until tensile wasand reached, which will the stability reduces significantly to 4.5 MPastate as soon as the working face advances beyond ofthe monitor points. significantly to 4.5 MPa as soon as the working face advances beyond the monitor points. The stresses entry retaining. The stresses in the gateway roof changed over the whole process from the beginning of the caving in the gateway roof changed over the whole process from the beginning of the caving to the roof to the roof stabilization, until astate tensile waswhich reached, will thethestability stabilization, until a tensile was state reached, will which influence theinfluence stability of gob-sideof the gob-side entry retaining. entry retaining.
Figure 5. Roof rock mass stress evolution law in working face end.
4. Force States of Roof Rock Mass Here, combining with the numerical results and field results, Figure 5. Roof rock mass stress evolution lawinvestigation in working face end. the force states of Figure 5. Roof rock mass stress evolution law in working face end. roof rock mass was analyzed to facilitate the failure mechanism analysis of roof rock mass and later 4. theoretical Force States of Roof Mass analysis of Rock supporting. 4. Force States of Roof Rock law Mass The stress evolution showed in Figure 5 experienced the front abutment pressure zone Here, combining with the numerical results and field investigation results, the force states of (section A-A in Figure 1) and stress decreasing zone (section B-B in Figure 1). In section A-A, there are Here, combining the to numerical results and field investigation results, the force states of roof rock mass was with analyzed facilitate the failure mechanism analysis of roof rock mass and later lateral forces F1a and F1b parallel to the rock strata and an overburden pressure P1 resembling the stress roof rock massanalysis was analyzed to facilitate the failure mechanism analysis of roof rock mass and later theoretical of supporting. The stress evolution law showed in Figure 55 experienced the front abutment pressure zone theoretical analysis of supporting. (section A-A evolution in Figure 1) law and stress decreasing zone5(section B-B in Figure 1). In abutment section A-A,pressure there are zone The stress showed in Figure experienced the front lateral forces F 1a and F1b parallel to the rock strata and an overburden pressure P1 resembling the stress (section A-A in Figure 1) and stress decreasing zone (section B-B in Figure 1). In section A-A, there
5 are lateral forces F1a and F1b parallel to the rock strata and an overburden pressure P1 resembling the
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stress Minerals (A1 , A2015, A3 ) in A Zone in Figure 5. In addition, field investigation showed that, there are 2 and 5, page–page bolt, mesh, and cable support force T1 , reinforced supporting force R1 in advance, shear forces Q1a , 1, A2 and A3) in A Zone in Figure 5. In addition, field investigation showed that, there are bolt, mesh, Q1b on(Athe two sides and gravity G. The force state was shown in Figure 6a. and cable support force T1, reinforced supporting force R1 in advance, shear forces Q1a, Q1b on the two In section B-B, the gob side lateral force disappears along the strata strike direction resembling sides and gravity G. The force state was shown in Figure 6a. the stress (A1 ) in B Zone in Figure 5, because of the roof rock mass caving near the gob side and part In section B-B, the gob side lateral force disappears along the strata strike direction resembling of the the roof rock(Amass fracturing, but the coal side lateral force is F2b resembling (A3 ) in B stress 1) in B Zone in Figure 5, because of the roof rock mass caving near the gobthe sidestress and part Zone in Figure 5. On the other hand, the overburden of pressure is assumed to become 0 resembling of the roof rock mass fracturing, but the coal side lateral force is F2b resembling the stress (A3) in B the stress in B Zone in Figure 5, due the roof separating maintoroof. There are also a bolt, Zone(A in2 )Figure 5. On the other hand, theto overburden of pressure from is assumed become 0 resembling stress (Asupport 2) in B Zone in Figure 5, due to the support roof separating from roof. There are end also and a bolt, mesh,the and cable force T2 , a strengthen force R the working face a shear 2 atmain mesh, and cable support force TG. 2, aThe strengthen support force R2 atinthe working face end and a shear force Q in field, also the gravity force state was shown Figure 6b. 2b force Q2b in field, therock gravity G. The forceinstate wassection shown in Figure 6b. Furthermore, thealso roof mass stress cross C-C (in Figure 1) has experienced the Furthermore, the roof rock mass stress in cross section C-C (in Figure 1) has experienced the stress states of sections A-A and B-B, but, field investigation showed that it will change after the stress states of sections A-A and B-B, but, field investigation showed that it will change after the construction of the artificial filling wall and the main roof rotary sinking, and the lateral force F on construction of the artificial filling wall and the main roof rotary sinking, and the lateral force Fa on a the coal the overburden pressure thelateral lateral force on the sidechange will change theside coal and side and the overburden pressurePPwill will restore, restore, the force on the gob gob side will to Fb . to InFaddition, the bolt, mesh and cable support force become to T, the shear forces at the two ends b. In addition, the bolt, mesh and cable support force become to T, the shear forces at the two ends become Q1a and 1b and gravityremain remain the the same force state waswas shown in Figure 6c. become to Q1atoand Q1bQand gravity sameG.G.The The force state shown in Figure 6c.
Figure 6. Roof rock mass force; (a) Force of cross section A-A; (b) Force of cross section B-B; (c) Force
Figure 6. Roof rock mass force; (a) Force of cross section A-A; (b) Force of cross section B-B; (c) Force of cross section C-C. of cross section C-C. 5. Mechanism for Failure of Roof Rock Mass
5. Mechanism for Failure of Roof Rock Mass
Roof failure was affected by many factors, so we will combine them with the failure characteristics
obtained by filed investigation, stress evolution law obtained numerical modeling thethe forcefailure Roof failure was affected by many factors, so we by will combine them and with states to study the roof failure mechanism. characteristics obtained by filed investigation, stress evolution law obtained by numerical modeling and the force states to study the roof failure mechanism. 5.1. Mechanism for Failure of Roof Rock Mass in Working Face End
5.1. Mechanism for Failure of Roof Rock Mass in Working Face End 5.1.1. Unloading Effect of the Lateral Stress
AffectedEffect by theofabutment pressure 5.1.1. Unloading the Lateral Stressin the mining and excavation process, the roof rock mass is in the yield state, as shown in cross section A-A in Figure 1. The pre-tensioned bolt-cable-mesh
Affected bysystem the abutment in significantly the mining increased and excavation the altered roof rock supporting improved pressure the strength, the yield process, strength, and the mass is in the yield state, as shown in cross section A-A in Figure 1. The pre-tensioned bolt-cable-mesh deformation characteristics of the roof rock mass. At the same time, the support system exerted a supporting system strength, significantly increased thehad yield strength, and altered pressure stress improved on the rock the mass, so the compressive zone stress state to be altered, which can the offset some of the tensile stress friction enhance the shear In support addition, the axial and deformation characteristics of theand roof rockand mass. At the samecapacity. time, the system exerted a lateral anchored forces shear strength of the weak structural thewhich roof can pressure stress on the rockincreased mass, sothethe compressive zone stress state plane, had topreventing be altered, rock mass from moving and sliding along the block structure plane. The pre-tensioned support offset some of the tensile stress and friction and enhance the shear capacity. In addition, the axial system controlled the expansion deformation and destruction, preventing roof separation, sliding, and lateral anchored forces increased the shear strength of the weak structural plane, preventing the fracture opening, and new crack generation in the anchorage zones, not only maintaining the roof rock massbut from sliding along the structure block structure plane. The [24,25]. pre-tensioned support integrity alsomoving forming and a pre-tensioned bearing with a large stiffness system controlled the expansion deformation and destruction, preventing roof separation, sliding, 6 fracture opening, and new crack generation in the anchorage zones, not only maintaining the integrity but also forming a pre-tensioned bearing structure with a large stiffness [24,25]. 712
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Influenced by the rotary sinking of the overlying roof and the superposition of the overburden pressure, the vertical displacement of the anchor bolt and cable increased significantly, leading to an increase of T1 . The stress increment of the roadway roof rock mass mainly comes from the bulking deformation. When the bolt, mesh, and cable apply pressure on the broken rock mass and adds an anchoring force, the rock mass volume or volume rise rate decrease, and a bulking force is produced by the broken rock mass that would work the bolts and surrounding rock at the same time and put the roof rock mass in an extrusion state. As shown in cross section A-A in Figure 1, it is precisely because of the interaction of the “support-surrounding rock”, that the roof rock mass presented a failure state in Figure 2a. Although the single hydraulic props support the roof strata with pressure, the extrusion state rock mass above the hydraulic support started loosening on the working face end. In cross section B-B in Figure 1, the loosening roof rock mass caved behind the support body on the gob side, causing a sharp reduction of the lateral pressure, especially for the accumulated bulking force. Meanwhile, the overburden pressure decreased almost to 0 due to the separation of the soft roof from the hard roof above. It can be observed that, accompanied by the gob roof caving, the lateral pressure unloaded and the bulking force decreased in the broken rock mass of the roadway roof, leading to its stress state changing from a three-dimensional stress state to a two-dimensional or uniaxial stress state, leading more easily to failure. At this point, together with the tensile stress effect, the rock mass volume “elastic” expansion in the layer strike direction caused the generation and propagation of a large number of cracks and the further failure of the rock mass. 5.1.2. Tensile and Shear Failure of the Working Face End Lower Layer Rock Mass Through the field investigation of the roof support body layout and its work process and based on the features of the roof rock mass deformation and failure, it is found that the working face end soft roof presents the following phenomenon: (1) An uneven supporting at the working face end roof. A strong mine ground pressure appeared, causing the significant sinking of the working face end roof in the working face advancing process, so a single hydraulic prop with an articulated roof beam was usually set as a reinforcement support. There is uneven pressure on the roof surface due to the low strength and stiffness of the sandy mudstone and mudstone and the higher strength and stiffness strengthening the support body, resulting in part of the support body inserting the roof and leading to a roof rock shear failure with shear stress q, as shown in Figure 7. In addition, the passive supporting force is too large and does not couple support with the bolts (cables), causing part of the bolts (cables) to be loosened, affecting the roof rock mass local stress state and resulting in ultimate rock mass failure. (2) Effect of local moment. Treating the reinforcement body (individual hydraulic prop and articulated roof beam, etc.) as the fulcrum, the lower roof strata produced a local bending on a small scale, but when the fulcrum bending moment (M) is too large, the layered roof strata cause tensile and compressive damage to the upper surface and lower surface, respectively, additionally increased the possibility of the support body inserting the upper roof and leading to rock mass shear failure, as shown in Figure 7. The two phenomena above are shown in the Figure 1 A-A and B-B cross sections and may occur simultaneously. 5.1.3. Dynamic Failure The working face end roof rock mass is often affected by vibration as described above. Here, 3DEC was also applied to study the dynamic failure characteristics of the roof rock mass, the coal and rock mass physical and mechanical parameters and the boundary conditions used in the 713
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numerical model shown in Figure 4, additionally adding the model viscous boundary conditions for all boundaries make the stress wave propagate or be absorbed to simulate the infinite foundation Minerals 2015,to 5, page–page environment. Minerals 2015,However, 5, page–page we alter the length to 1.6 m in the y direction and treat it as a plane strain environment. However, we alter the length m in 1, theunder y direction and treat as a planestress strainwave model. As the location in cross section A-A to in 1.6 Figure the action of ait triangle Minerals 2015, 5, page–page model. As the location in cross section A-A in Figure 1, under the action of a triangle stress wave to environment. However,ofwe alter the length to 1.6 mtointhe theroof y direction and treat it assimulation a plane strain to simulate the influence a dynamic disturbance rock mass. In the process, simulate the influence of a dynamic disturbance to the roof rock mass. In the simulation process, the model. As the location inwe cross section A-Awave intoFigure the action atreat triangle stress wave to the stress peak value of the triangle stress is m 8 1, MPa rise time isand 1ofms and time is environment. However, alter the length inunder the ytime direction it the as a decrease plane strain stress peak value of the triangle stress wave is1.6 8 the MPa rise is 1 In msthe and the decrease time is simulate the influence of a dynamic disturbance to roof rock mass. simulation process, the As the location in cross section A-A in Figure 1, under the action of a triangle stress wave to 7 ms 7model. [26–29]. Roof bolt (like a , a , a , a , a , a in Figure 4c) end displacement and the surrounding 1a1, a22, a33, a4, 4a5, a56 in6Figure 4c) end displacement and the surrounding rock ms [26–29]. Roof bolt (like stress peak of the stress wave is to 8 MPa riserock timemass. is 1 ms and the decrease timethe is simulate thevalue influence of triangle a dynamic theshown roof In9,8the simulation process, rock plastic zone weremonitored, monitored, the results are in Figures and 9 respectively. plastic zone were andand thedisturbance results are shown in Figures 8 and respectively. 7stress ms [26–29]. Roof bolt (like a1, a2, a3stress , a4, a5,wave a6 in Figure 4c) end rock peak value of the triangle is 8 MPa risedisplacement time is 1 ms and andthe thesurrounding decrease time is plastic zone were monitored, and the results are shown in Figures 8 and 9, respectively. 7 ms [26–29]. Roof bolt (like a1, a2, a3, a4, a5, a6 in Figure 4c) end displacement and the surrounding rock plastic zone were monitored, and the results are shown in Figures 8 and 9, respectively.
Figure 7. Schematic of support body damage to the roof rock mass.
Figure 7. Schematic of support body damage to the roof rock mass. Figure 7. Schematic of support body damage to the roof rock mass. Figure 7. Schematic of support body damage to the roof rock mass.
Figure 8. Displacement variation before and after the dynamic disturbance. Figure 8. Displacement variation before and after the dynamic disturbance.
Figure 8. Displacement variation before and after the dynamic disturbance. Figure 8. Displacement variation before and after the dynamic disturbance.
Figure 9. Plastic zone of surrounding rock: (a) Plastic zone before disturbance; (b) Plastic zone after disturbance. Figure 9. Plastic zone of surrounding rock: (a) Plastic zone before disturbance; (b) Plastic zone after disturbance. Figure 9. observed Plastic zone surrounding rock: (a) (b) Plastic zone It can be from Figure 8 that the displacement ofbefore all thedisturbance; monitoring points had a small Figure 9. Plastic zone of ofsurrounding rock: (a) Plastic Plasticzone zone before disturbance; (b) Plastic zone after disturbance. increase in magnitude after the disturbance.
afterItdisturbance. can be observed from Figure 8 that the displacement of all the monitoring points had a small Figure 9a,b show after the plastic distribution zone before and after the dynamic disturbance of the increase in magnitude the disturbance. It can be observed from Figure 8 that theThe displacement of allcountless the monitoring points a small gateway surrounding rock mass, respectively. stress wave had reflections andhad refractions Figure 9a,b showfrom the plastic distribution zone before and the monitoring dynamic disturbance of the It can beinobserved Figure 8 that the displacement ofafter all the points had a small increase magnitude after the disturbance. on the roof crack surface according to the stress wave propagation theory, and the tensile stress would gateway surrounding rock mass, respectively. The stress wave had countless reflections and refractions Figure 9a,b show thethe plastic distribution zone before and after the dynamic disturbance of the increase in magnitude after disturbance. on the roof crack surface according to the stressThe wave and the tensile stress would 8 propagation gateway surrounding rock mass, respectively. stress wave hadtheory, countless reflections and refractions on the roof crack surface according to the stress wave 8 propagation theory, and the tensile stress would
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Figure 9a,b show the plastic distribution zone before and after the dynamic disturbance of the gateway surrounding rock mass, respectively. The stress wave had countless reflections and refractions on the roof crack surface according to the stress wave propagation theory, and the tensile stress would cause changes to the roof rock mass stress state and result in a combination Minerals 2015, 5,on page–page of shear failure existing joints/weakness horizons, an extension of critically oriented joints and propagation of new fractures through previously intact rock so that the integrity of the roof became cause changes to the roof rock mass stress state and result in a combination of shear failure on existing poor, the plastic zone expanding into the higher strata, and the ultimate widespread destruction of joints/weakness horizons, an extension of critically oriented joints and propagation of new fractures the roof rock mass [7]. At the same time, the two gateway sides and floor rock mass also produced through previously intact rock so that the integrity of the roof became poor, the plastic zone expanding plastic damage under theand stress wave action, but had small extended failure zones. into the higher strata, the ultimate widespread destruction of the roof rock mass [7]. At the same The above results show that the increment of the displacement and plastic was small, time, the two gateway sides and floor rock mass also produced plastic damage underzone the stress wave but due to the but constant increase of the dynamic action, had small extended failure zones. load, it can be presumed that the working face end results show the increment of the displacement and plastic zoneunder was small, but soft roof The rockabove deformation andthat failure magnitude would significantly increase the repeated due to the action, constantand increase of the dynamic load, it can be presumed that the caused workingtoface softmass. dynamic load the more cracks are formed, the more damage theend rock roof rockthe deformation and failure would significantly under repeated To minimize power damage, theremagnitude is a need to improve the soft increase roof rock massthe stress state and dynamic load action, and the more cracks are formed, the more damage caused to the rock mass. reduce the development degree of cracks. To minimize the power damage, there is a need to improve the soft roof rock mass stress state and reduce the development cracks. 5.2. Mechanism for Failure of degree the RoofofRock Mass in Retaining Roadway
As in Figure 2c,ofcross section C-C in stabilized zone in retaining roadway experience 5.2. shown Mechanism for Failure the Roof Rock Mass in the Retaining Roadway a fractureAs deformation process at the working face end that can be in influenced by the main roof rotary shown in Figure 2c, cross section C-C in the stabilized zone retaining roadway experience deformation stress recovery process besideface theend working resultingbyinthe itsmain larger damage a fractureand deformation process at the working that canface, be influenced roof rotary and deformation degree, causing a weakening of its integrity. After support ofand the roof deformation and stress recovery process beside the working face,strengthening resulting in its the larger damage rock mass in the retaining roadway and constructing the artificial filling wallthe body, theirofintegrity deformation degree, causing a weakening of its integrity. After strengthening support the roof and rockwere massimproved in the retaining roadway and constructing the artificial filling wall body, integrity stability and the bulking force was restored to a certain degree. For atheir certain retaining and stability improved and the bulking forcerecycled, was restored certain degree. For wall’s a certain distance, after thewere strengthened support body was due to to athe artificial filling higher retaining distance, after the strengthened support body was recycled, due to the artificial filling wall’s strength and stiffness, the bulking force would release along its inside. Therefore, the soft roof of higher strength and stiffness, the bulking force would release along its inside. Therefore, the soft roof retained entry caved at the man-made constructed wall (see shear failure line in Figure 10) due to of retained entry caved at the man-made constructed wall (see shear failure line in Figure 10) due to forces such as the overburden pressure (P), the bulking stress (Ps ), rock mass gravity (G) and the bolt, forces such as the overburden pressure (P), the bulking stress (Ps), rock mass gravity (G) and the bolt, mesh,mesh, and and cable support force formeda a“similar “similar cantilever beam” structure, as in shown cable support force(T), (T), which which formed cantilever beam” structure, as shown in Figure 10. Figure 10.
Figure 10. Roof rock mass stress model of the retaining roadway.
Figure 10. Roof rock mass stress model of the retaining roadway.
6. Roof Support Countermeasures
6. Roof Support Countermeasures
To develop a supporting method for the roof rock mass at the working face end and retained
To develop supporting method massbe atanalyzed. the working facebolt endis and retained gateway, the a deformation and force infor the the roofroof rock rock mass must The rock the basic supporting tool. First, we calculated “cable” element to represent rock boltsThe on the working gateway, the deformation and force inthe the roof rock mass must be the analyzed. rock bolt is the end withtool. several different bolt installation angleselement in 3DEC to find the bolt basicface supporting First, we calculated the “cable” to model represent the rock boltsend on the displacement change characteristics. Then, we analyzed the bolt limit equilibrium tension force working face end with several different bolt installation angles in 3DEC model to find the bolt end change features in the retained gateway roof rock mass with the change of the bolt installation angle though the force balance equation.
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displacement change characteristics. Then, we analyzed the bolt limit equilibrium tension force change features in the retained gateway roof rock mass with the change of the bolt installation angle Minerals 2015, 5, page–page though the force balance equation. 6.1. Deformation Analysis of Working Face End Roof
characteristics, the following mechanism for the From the soft roof failure process and failure characteristics, obtained: soft roof support can be obtained: (1) Providing the roof rock mass extrusion stress and changing the rock mass stress state to improve their Preventing cracks from generating andand propagating to increase both their strength strength[11]; [11];(2)(2) Preventing cracks from generating propagating to increase both the strength the affected roof strata andstiffness the stiffness the whole bolted to reduce the strength of theofaffected roof strata and the of theofwhole bolted stratastrata to reduce roof roof deformation dynamic damage; Reducing brokenrock rockmass massbulking bulking force force and deformation and and dynamic damage; andand (3) (3) Reducing thethebroken maintaining the stability of the roadway roof roof by by supporting supporting in in time. time. According to the above points and the “bolt extruding reinforcement theory” by pre-tensioned bolts providing a compressive zone in the axial direction [11], the working face end soft roof mechanical structure should be similar to ˝< < α
