sustainability

Article Analysis and Optimization of Entry Stability in Underground Longwall Mining

Yubing Gao 1,2, Dongqiao Liu 1,2, Xingyu Zhang 1,2 and Manchao He 1,* 1 State Key Laboratory of Geomechanics & Deep Underground Engineering, China University of Mining & Technology, Beijing 100083, China; [email protected] (Y.G.); [email protected] (D.L.); [email protected] (X.Z.) 2 School of Mechanics and Civil Engineering, China University of Mining & Technology, Beijing 100083, China * Correspondence: [email protected]

Received: 1 October 2017; Accepted: 10 November 2017; Published: 12 November 2017

Abstract: For sustainable utilization of limited coal resources, it is important to increase the coal recovery rate and reduce mine accidents, especially those occurring in the entry (gateroad). Entry stabilities are vital for ventilation, transportation and other essential services in underground coal mining. In the present study, a finite difference model was built to investigate stress evolutions around the entry, and true triaxial tests were carried out at the laboratory to explore entry wall stabilities under different mining conditions. The modeling and experimental results indicated that a wide coal pillar was favorable for entry stabilities, but oversize pillars caused a serious waste of coal resources. As the width of the entry wall decreased, the integrated vertical stress, induced by two adjacent mining panels, coupled with each other and experienced an increase on the entry wall, which inevitably weakened the stability of the entry. Therefore, mining with coal pillars always involves a tradeoff between economy and safety. To address this problem, an innovative non-pillar mining technique by optimizing the entry surrounding structures was proposed. Numerical simulation showed that the deformation of the entry roof decreased by approximately 66% after adopting the new approach, compared with that using the conventional mining method. Field monitoring indicated that the stress condition of the entry was significantly improved and the average roof pressure decreased by appropriately 60.33% after adopting the new technique. This work provides an economical and effective approach to achieve sustainable exploitation of underground coal resources.

Keywords: sustainable exploitation; coal resources; longwall mining; entry stability

1. Introduction Coal is among the most important foundational energies, accounting for approximately 30% of the total energy consumption around the world [1,2]. For sustainable utilization of limited coal resources, it is imperative to increase the coal recovery rate and improve mining safety. Failure or collapse of the entry, which can lead to human loss or interruption of mining, is a common accident in the mine. Entry stabilities are related to the surrounding stress conditions. Mahdi et al. [3] carried out a numerical investigation on how longwall mining affected the stress state in the border area and how the stress changes would affect future mining. They found that the pillar stress fluctuated up and down during mining, and the failure zone of a pillar was approximately 12 m from the entry rib to the wall center. Suchowerska et al. [4] investigated some variables that affected stress redistributions. They found that the abutment angle of the pillar loading has a greater effect than other factors such as the overburden depth. Basarir et al. [5] predicted the stresses around the entry using empirical and numerical analysis of different mining methods (the fully mechanized traditional longwall and longwall top coal caving methods). Such issues have also been addressed in the literature [6–10]. These studies showed that the stress redistribution around the entry is associated with many factors such as geologic setting,

Sustainability 2017, 9, 2079; doi:10.3390/su9112079 www.mdpi.com/journal/sustainability Sustainability 2017, 9, 2079 2 of 19 working face layout and mining disturbances. Certainly, the entry surroundings, especially the coal pillar between two adjacent mining panels, has an important influence on stress distributions around the entry. In terms of entry wall (coal pillar) stabilities, most previous studies were conducted through theoretical analysis and numerical simulation. For example, Gao [11] evaluated pillar stabilities using the limiting-equilibrium theory. In his study, the coal pillar was divided into three zones: the loose zone, plastic zone and elastic zone. Wang et al. [12] proposed an approach to simulate the static and dynamic behaviors of the coal pillar. They concluded that the width-to-height (W/H) ratio of the coal pillar was important in affecting entry wall stabilities. Bieniawski, Reed and other researchers [13–15] developed different methods to design and optimize the pillar size. Presently, gob-side entry retaining by filling artificial materials is a frequently used non-pillar mining method to optimize sustainable mining. In this method, the former entry is artificially retained as the tailgate for the next mining panel by using pigsties, paste-like backfill material, high-water packing material, and other fill materials. In this mining pattern, non-pillar is left after mining. However, excessive consumption of filling materials and high stress concentration on the filling materials set limits on the popularity of this method applied in complicated geological conditions, such as thick coal seams. Therefore, it is necessary to find a more effective and economical method to improve mining sustainability and safety. Previous studies reveal that the entry stability is primarily determined by its surrounding stresses [16–18]. Considering this, we proposed a stress optimization method that aims to increase the coal recovery rate and improve entry stabilities. First, the stability of the entry wall in conventional mining patterns was systematically studied using numerical modeling, laboratory test and theoretical analysis. Subsequently, a novel roof addressing technique was proposed and tested in the field. The applicative effects indicate that the proposed approach is effective and has broad application prospects in sustainable mining of underground coal resources.

2. Analysis of the Entry Stability in Conventional Mining Patterns Technological advances in mining mechanization over the past decades have enabled longwall mining to supersede room and pillar mining in terms of both safety and production levels [19]. The dimensional design of the chain pillar between two mining panels is one of the main focus points in longwall mining. Entry wall stabilities to a great extent determine entry stabilities, especially in a high-stress entry. To explore the vulnerable surroundings of the entry and the reason that causes the entry to be unstable, numerical simulation, laboratory test and theoretical analysis were conducted, respectively. Since the width of the coal pillar is an important factor that influences stress distributions and entry stability, several scenarios that emphasizes the pillar width were performed. Distribution laws of the vertical stress around the entry were studied using numerical simulation, and failure states of the coal pillar were investigated using laboratory tests. Based on the obtained results, we then proposed pertinent strategies to address the instability problem of the entry.

2.1. Numerical Analysis of the Entry Wall Stability

2.1.1. Study Site The Ningtiaota coal mine is located in Shanxi province, China. This mine produces approximately 20 million tons of coal annually. The selected panels are approximately 200~300 m wide and 3000 m long, with a dip angle of less than 2◦. The mean thickness of the coal seam is 4.0 m, with an average overburden varying from 80 to 160 m. Figure1 shows the mine location and column map of the stratum. The roof strata are mainly composed of hard sandstones, and the floor strata consist of siltstone (8.1 m) and fine-grain sandstone (12.3 m). Full-seam, comprehensive mechanized and retreating mining methods were used in the panels, and a spontaneous caving method was used to treat the gob roof during the mining process. Sustainability 2017, 9, 2079 3 of 19

mechanized and retreating mining methods were used in the panels, and a spontaneous caving method was used to treat the gob roof during the mining process.

Sustainability 2017, 9, 2079 3 of 19

Sustainabilitymechanized2017, 9and, 2079 retreating mining methods were used in the panels, and a spontaneous caving3 of 19 method was used to treat the gob roof during the mining process.

Figure 1. Location of the mine and lithology of the rock layers.

2.1.2. Numerical Model To investigate evolution laws of the vertical stress on the entry wall under different mining conditions, a 3D numerical model was developed using the finite difference method, as illustrated in Figure 2. The dimensionsFigureFigure 1. of 1.Location theLocation model of of the thewere mine mine a length and and lithologylithology of 300 ofm, the the width rock rock oflayers. layers. 200 m, and height of 150 m, based on the geological column. In this study, the excavation was performed by changing the 2.1.2. Numerical Model 2.1.2.mechanical Numerical properties Model of the coal seam in the caved area to a very soft elastic material, in which case, the bearingTo investigate capacities evolution of the cavedlaws ofrock the could vertical be stressfully considered.on the entry This wall analogy under different method miningof the To investigate evolution laws of the vertical stress on the entry wall under different mining excavationconditions, process a 3D numerical has also been model used was by developed other researchers using the [20,21]. finite difference method, as illustrated conditions, a 3D numerical model was developed using the finite difference method, as illustrated in FigureIn terms 2. The of dimensions the boundary of the constraints, model were the a length left and of 300 right m, widthborders of 200restrict m, and the height horizontal of 150 in Figuredisplacement,m, based2. Theon the dimensionsand geological the lower of column. the bounder model In this wererestricts study, a length thethe excavationvertical of 300 m,displacement, was width performed of 200 respectively. m,by changing and height The the of 150minimummechanical m, based horizontal on properties the geological stress of the‐to coal‐vertical column. seam stress in In the this ratio caved study, is areakept the to as a excavation 0.8very according soft elastic was to performed material,the in‐situ in stresses bywhich changing case, of theNo.the mechanical 2bearing‐2 coal capacities propertiesseam. The of ofconstitutive the the caved coal seam equationrock incould the adopted cavedbe fully areain considered.this to model a very Thisis soft Mohr–Coulomb elasticanalogy material, method criterion, inof which the case,whichexcavation the bearingis a mathematical process capacities has also model of been the describing caved used by rock otherthe could response researchers be fully of brittle [20,21]. considered. materials This to shear analogy stress method as well ofas the excavationnormalIn stress. processterms of has the also boundary been used constraints, by other researchersthe left and [20 right,21]. borders restrict the horizontal displacement, and the lower bounder restricts the vertical displacement, respectively. The minimum horizontal stress‐to‐vertical stress ratio is kept as 0.8 according to the in‐situ stresses of No. 2‐2 coal seam. The constitutive equation adopted in this model is Mohr–Coulomb criterion, which is a mathematical model describing the response of brittle materials to shear stress as well as normal stress.

FigureFigure 2. 2.Schematic Schematic diagram of of the the numerical numerical model. model.

In termsThe properties of the boundary of real rock constraints, masses thein the left field and right are usually borders different restrict the from horizontal those tested displacement, at the andlaboratory the lower [22]. bounder The elastic restricts modulus, the vertical cohesion displacement, and tensile strength respectively. results are The generally minimum 0.1~0.25 horizontal of the laboratory testing results. The Poisson ratios are approximately 1.2~1.4 of the laboratory testing stress-to-vertical stress ratio is kept as 0.8 according to the in-situ stresses of No. 2-2 coal seam. The constitutive equation adoptedFigure 2. in Schematic this model diagram is Mohr–Coulomb of the numerical criterion, model. which is a mathematical model describing the response of brittle materials to shear stress as well as normal stress. The properties of real rock masses in the field are usually different from those tested at the laboratoryThe properties [22]. The of elastic real rock modulus, masses cohesion in the fieldand tensile are usually strength different results are from generally those tested0.1~0.25 at of the laboratorythe laboratory [22]. The testing elastic results. modulus, The Poisson cohesion ratios and are tensile approximately strength 1.2~1.4 results of are the generally laboratory 0.1~0.25 testing of the laboratory testing results. The Poisson ratios are approximately 1.2~1.4 of the laboratory testing results. On the basis of the laboratory test results of the mechanical properties of the rock samples, the final property parameters of the rock strata in the numerical model are presented in Table1. Sustainability 2017, 9, 2079 4 of 19

Table 1. Mechanical parameters of the rock strata.

Density Elasticity Friction Angle Cohesion Tensile Strength Rock Strata (kg/m3) Modulus (GPa) (deg) (MPa) (MPa) Overlying strata 2500 15.2 30 1.8 0.7 Fine sandstone 2650 19.6 32 2.3 1.2 Medium sandstone 2600 13.5 34 2.4 1.4 Quartz sandstone 2780 21.1 38 3.2 1.6 Siltstone 2550 16.5 30 2.1 0.9 No. 2-2 coal seam 1250 4.3 19 0.9 0.4 Siltstone 2550 13.9 26 3.0 1.2 Fine sandstone 2650 18.8 33 3.1 1.5

2.1.3. Evolution Laws of the Vertical Stress on the Entry Wall In the numerical modeling, plane I and plane I were successively exploited to illustrate and analyze the recorded stress on the entry wall. The widths of the entry wall between planes I and II were 80 m, 50 m, 30 m, 15 m, 8 m and 5 m, respectively, which can be used to represent different mining patterns. The contour maps of the vertical stress are presented in Figure3. According to the simulation results, the following can be concluded:

(1). In terms of stress migration, the stress peaks moved closer to each other as the pillar width decreased. The macroscopic shape of the vertical stress was double-arched when the pillar widths were large (80 m, 50 m and 30 m). Gradually, the double-arched shape of the stress evolved into a single-arched shape (15 m, 8 m and 5 m), and the integrated peak stress increased as well. In fact, the vertical stress on the pillar was induced in two stages. In the first stage, the stress was induced by the production of the present panel (panel I). After the surrounding rocks had stabilized, the mining-induced stress gradually reached equilibrium and remained constant. Subsequently, as panel II was mined, and an additional stress was then superimposed on the existing stress. Finally, the vertical stress achieved equilibrium on the entry wall. (2). In terms of the stress magnitudes, the position of the stress peak was in front of the panel. When the pillar width was more than 50 m, two stress peaks caused by panel I and II were distributed near the pillar sides. As the pillar width decreased, the central stress increased on the pillar, whereas the peak value remained unchanged (approximately 21.8 MPa). When the pillar widths were from 50 m to 30 m, there were also two stress peaks distributed on the two sides. However, the peak values increased quickly with decreasing pillar width. When the pillar width was less than 15 m, there was only one stress peak indicated macroscopically. The maximum vertical stress reached 27.8 MPa when the pillar width was 15 m. Interestingly, as the pillar width continued to decrease to 5 m, the peak stress decreased, as shown in Figure3f. In a narrow pillar, there would be more mining-induced fractures distributed in the coal body. Essentially, the entry wall had been damaged. These fractures provided voids for the stress to release, causing the vertical stress to decrease.

As seen in the simulation results, there were always concentrated stresses distributed on the entry wall, regardless of the pillar size. Specifically, when the pillar was wide, the concentrated stress was relatively small. As the pillar width decreased, the overall vertical stress increased. When the pillar was too narrow to bear the concentrated stress, the entry wall coal was damaged. Accordingly, we can conclude that the concentrated stress was the main reason accounting for entry instabilities. Sustainability 2017, 9, 2079 5 of 19 Sustainability 2017, 9, 2079 5 of 19

FigureFigure 3. Evolution3. Evolution laws laws of the of verticalthe vertical stress stress on the on entry the entry wall. wall. (a) The (a) pillar The widthpillar width was 80 was m. (80b) Them. ( pillarb) widthThe was pillar 50 width m. (c) was The 50 pillar m. width(c) The was pillar 30 width m. (d) was The 30 pillar m. width(d) The was pillar 15 m.width (e) Thewas pillar15 m. width(e) The was 8 m.pillar (f) The width pillar was width 8 m. ( wasf) The 5 m.pillar width was 5 m.

2.2.2.2. Experimental Experimental Analysis Analysis of of the the Entry Entry Wall Wall Stability Stability

2.2.1.2.2.1. Model Model Description Description at at the the Laboratory Laboratory AccordingAccording to to rock rock damage damage theory,theory, highhigh stresses on on the the entry entry wall wall tend tend to toproduce produce more more mining‐induced fractures [23–25]. The degree of damage differs as the pillar width changes. mining-induced fractures [23–25]. The degree of damage differs as the pillar width changes. Simulation results indicated that when the pillar was wide, the stress peaks were distributed at the Simulation results indicated that when the pillar was wide, the stress peaks were distributed at sides. In this case, most fractures were distributed at the sides, and the central zone remained the sides. In this case, most fractures were distributed at the sides, and the central zone remained almost undamaged. Therefore, the central cubic zone was restricted in the horizontal direction and almostcould undamaged. be simplified Therefore, to be laterally the confined, central cubic as shown zone in was Figure restricted 4a. in the horizontal direction and could beIn simplified contrast, when to be the laterally entry confined,wall was asnarrow, shown the in global Figure 4stressa. was large, thus causing more voidsIn contrast, and mining when‐induced the entry fractures. wall was In this narrow, case, thethe globalhorizontal stress limit was was large, weakened, thus causing and the more lateral voids andload mining-induced accordingly reduced, fractures. as Inshown this case, in Figure the horizontal 4b,c. Combining limit was these weakened, results with and the lateralpractical load accordingly reduced, as shown in Figure4b,c. Combining these results with the practical situation on Sustainability 2017, 9, 2079 6 of 19 Sustainability 2017, 9, 2079 6 of 19 Sustainability 2017, 9, 2079 6 of 19 site,situation we used on cubic site, coalwe used units cubic confined coal withunits differentconfined amountswith different of lateral amounts pressure of lateral to simply pressure show to the failuresimply behaviors show ofthe the failure entry wallbehaviors with differentof the widths.entry wall This with type ofdifferent simplification widths. has This also type been of used bysimplification other researchers has also [26– been28]. used by other researchers [26–28].

(a) (b) (c) (a) (b)K2γH (c) K2γH K γH K3γH KIγH 3 KIγH

γH I panel gob γH II panel gob I panel gob II panel gob I panel gob II panel gob I panel gob II panel gob I panel gob γH II panel gob I panel gob II panel gob γH γH γH

σ3 σ3 σ3 σ3 σ3 σ3 z σ σ σ3 σ σ3 σ3 z 3 3 3 x x σ σ σ2 y σ2 2 y σ2 2 σ2

FigureFigure 4. Model4. Model simplification simplification at at the the laboratory. laboratory. (a )(a A) wideA wide coal coal pillar. pillar. (b )( Ab) mediumA medium wide wide coal coal pillar. (c)pillar. A narrow (c) A coal narrow pillar. coal pillar.

2.2.2.2.2.2. Methodology Methodology The natural coal obtained on site was processed into cubic coal samples with a side length of TheThe natural natural coal coal obtained obtained on on site site waswas processed into into cubic cubic coal coal samples samples with with a side a side length length of of 110 mm. Three intact samples were selected to investigate the pillar failure behaviors. The average 110 mm. Three intact samples were selected3 to investigate the pillar failure behaviors. The average density of the samples was 1.25 g/cm3. These samples were tested in a true triaxial test unit, as densitydensity of theof the samples samples was was 1.25 1.25 g/cm g/cm3. These. These samples samples were were tested tested in ain true a true triaxial triaxial test test unit, unit, as shown as shown in Figure 5. This equipment is mainly composed of a loading system, a pressure pump, a in Figureshown5 .in This Figure equipment 5. This equipment is mainly composed is mainly ofcomposed a loading of system, a loading a pressure system, pump, a pressure a data pump, collection a data collection system and a triaxial cell. These specimens were first put into a metal box to ensure system and a triaxial cell. These specimens were first put into a metal box to ensure stability and then stability and then installed in the triaxial cell. The equipment can accomplish automatic installed in the triaxial cell. The equipment can accomplish automatic self-compensation of the pressure self‐compensation of the pressure and control the confining pressure by utilizing a rock servo and control the confining pressure by utilizing a rock servo controlling rheological testing machine. controlling rheological testing machine. This equipment can conduct the uniaxial compression test, Thisthe equipment triaxial conventional can conduct compression the uniaxial test compression and the cyclic test, the loading triaxial test, conventional etc. During compression the test, the test andexperimental the cyclic loading data of test, the stress etc. During‐strain curve the test, can the be experimentalrecorded by the data data of acquisition the stress-strain system curve utilizing can be recordedthe corresponding by the data sensors. acquisition system utilizing the corresponding sensors.

Figure 5. Test equipment at the laboratory. Figure 5. Test equipment at the laboratory. Based on the crustal stress of the test site, these samples were confined with different lateral pressuresBased on to thesimulate crustal different stress sized of the pillars, test site, as described these samples in Table were 2. Because confined the withstrike different length of lateralthe pressurespressures to to simulate simulate different different sized sized pillars, pillars, as described in in Table Table 2.2. Because Because the the strike strike length length of ofthe the pillar is much larger than the width, thus we assume that the lateral pressures in the x direction are pillar is much larger than the width, thus we assume that the lateral pressures in the x direction are all the same (6.0 MPa). The first sample (S1) was confined with 2.0 MPa in the y direction to all the same (6.0 MPa). The first sample (S1) was confined with 2.0 MPa in the y direction to simulate a wide coal pillar. The second sample (S2) was confined with 1.0 MPa in the y direction to simulate Sustainability 2017, 9, 2079 7 of 19

Sustainability 2017, 9, 2079 7 of 19 a medium sized coal pillar, and the third sample (S3) was confined with 0.5 MPa in the y direction to simulate a wide coal pillar. The second sample (S2) was confined with 1.0 MPa in the y direction to simulate a narrow coal pillar. First, a very small force was applied on the surface of the cubic samples simulate a medium sized coal pillar, and the third sample (S3) was confined with 0.5 MPa in the y at adirection rate of 0.5 to simulate kN/s. Subsequently, a narrow coalσ pillar.x and σFirst,y were a very increased small force to the was predetermined applied on the confining surface of stresses, the as listed in the Table2. Finally, the samples were loaded at a stable rate of 0.004 mm/s in the z direction cubic samples at a rate of 0.5 kN/s. Subsequently, σx and σy were increased to the predetermined untilconfining the maximum stresses, displacement as listed in the limitations Table 2. Finally, were reached.the samples were loaded at a stable rate of 0.004 mm/s in the z direction until the maximum displacement limitations were reached. Table 2. Confining stresses loaded on the samples. Table 2. Confining stresses loaded on the samples. Sample σz (Sample Faces) σx (Sample Faces) σ2 (MPa) σy (Sample Faces) σ3 (MPa) SampleS1 σz (Sample A & C Faces) σx (E(Sample & F) Faces) 6.0σ2 (MPa) σy (B(Sample & D) Faces) 2.0σ3 (MPa) S1 S2A && C C (E &(E F) & F) 6.06.0 (B &(B D) & D) 1.0 2.0 S3 A & C (E & F) 6.0 (B & D) 0.5 S2 A & C (E & F) 6.0 (B & D) 1.0 S3 A & C (E & F) 6.0 (B & D) 0.5 2.2.3. Test Results 2.2.3. Test Results Figure6 shows the damage states of each sample after testing. It is obvious that when the σy was

2.0 MPa,Figure the six 6 shows faces ofthe the damage S1 sample states almost of each remained sample after intact. testing. There It is were obvious only that some when micro-pores the σy or micro-fractureswas 2.0 MPa, the distributed six faces onof thethe surfaces,S1 sample indicating almost remained that the central intact. zoneThere of were a wide only pillar some is less likelymicro to‐ bepores damaged. or micro As‐fractures the pillar distributed width decreased, on the surfaces, the confining indicating stress that also the decreased.central zone As of seena wide from thepillar damage is less state likely of to S2 be sample, damaged. two As great the fracturespillar width and decreased, many pores the confining were generated stress also on decreased. the surfaces. Furthermore,As seen from when the damage the σy decreasedstate of S2 sample, to 0.5 MPa, two great the sample fractures was and crushed many pores into were blocks. generated The damage on morphologiesthe surfaces. of Furthermore, the S1, S2 and when S3 samples the σy decreased showed thatto 0.5 as MPa, the confining the sample stress was decreased, crushed into the blocks. degree of damageThe damage of the samples morphologies became of more the S1, serious. S2 and S3 samples showed that as the confining stress decreased,Figure7 showsthe degree the of stress-strain damage of curvesthe samples during became the test. more The serious. critical failure stresses of S1, S2 and Figure 7 shows the stress‐strain curves during the test. The critical failure stresses of S1, S2 and S3 were 35.397 MPa, 23.379 MPa and 16.591 MPa, respectively. As the confining stress decreased, S3 were 35.397 MPa, 23.379 MPa and 16.591 MPa, respectively. As the confining stress decreased, the peak stress decreased as well. From the post-peak behavior of the curves, we can see that when the peak stress decreased as well. From the post‐peak behavior of the curves, we can see that when the confining stress is large, the change of the loading stress is gentler, indicating that a strong lateral the confining stress is large, the change of the loading stress is gentler, indicating that a strong restriction caused the wide coal pillar to be more stable. lateral restriction caused the wide coal pillar to be more stable. TheThe damage damage morphologies morphologies and and stress-strain stress‐strain curves curves showed showed that that when when the the pillar pillar is wide, is wide, it would it bewould favorable be tofavorable avoiding to entry avoiding instability entry problems. instability However, problems. oversize However, pillars oversize cause pillars a serious cause waste a of resources.serious waste Although of resources. mining with Although narrow mining pillars with can narrow save more pillars resources, can save high more stress resources, concentrations high andstress less lateralconcentrations restrictions and causedless lateral the entryrestrictions wall to caused be more the unstable.entry wall Hence, to be more mining unstable. with coal Hence, pillars alwaysmining involves with coal a tradeoff pillars always between involves economy a tradeoff and safety. between economy and safety.

(a)

(b)

Figure 6. Cont.

Sustainability 2017, 9, 2079 8 of 19 Sustainability 2017, 9, 2079 8 of 19

Sustainability 2017, 9, 2079 8 of 19

(c) (c) FigureFigure 6. Damage6. Damage states states of theof the samples samples under under different different confining confining pressures. pressures. (a) Six(a) facesSix faces of S1 of sample. S1 Figure 6. Damage states of the samples under different confining pressures. (a) Six faces of S1 sample. (b) Six faces of S2 sample. (c) Six faces of S3 sample. (b) Six facessample. of S2(b) sample.Six faces of (c S2) Six sample. faces (c of) Six S3 faces sample. of S3 sample.

Figure 7. Z‐direction stress‐strain curves under different confining pressures. (a) S1 sample. (b) S2 Figure 7. Z-direction stress-strain curves under different confining pressures. (a) S1 sample. Figuresample. 7. Z‐direction (c) S3 sample. stress ‐strain curves under different confining pressures. (a) S1 sample. (b) S2 (b)sample. S2 sample. (c) S3 (c sample.) S3 sample. 2.3. Theoretically Analysis of the Coal Pillar Stability 2.3. Theoretically Analysis of the Coal Pillar Stability 2.3. TheoreticallyNumerical Analysis simulation of the Coal indicated Pillar Stabilitythat entry stability was highly related to its surrounding NumericalNumericalstresses. As simulation thesimulation pillar width indicated indicated decreased, that that entry the entry overall stability stability vertical was stresswas highly actinghighly related on related the to coal its topillar surrounding its increased.surrounding stresses. When the pillar was too narrow to bear the concentrated stress, the entry wall was damaged. The Asstresses. the pillar As width the pillar decreased, width decreased, the overall the vertical overall stress vertical acting stress on acting the coal on pillarthe coal increased. pillar increased. When the Whenlaboratory the pillar test was also too showed narrow that to as bear the pillarthe concentrated width decreased stress, (or confining the entry pressure wall was decreased), damaged. the The pillar wascoal too body narrow tended to to bearbe more the unstable. concentrated Hence, we stress, can understand the entry that wall the was width damaged. of the coal The pillar laboratory is laboratory test also showed that as the pillar width decreased (or confining pressure decreased), the test alsoan showed important that factor as thethat pillaraffects widththe entry decreased stability. In (or this confining section, the pressure entry wall decreased), stability with the regard coal body coal body tended to be more unstable. Hence, we can understand that the width of the coal pillar is tended toto bepillar more width unstable. was theoretically Hence, weinvestigated can understand in three frequently that the widthused mining of the patterns, coal pillar i.e., wide is an coal important an important factor that affects the entry stability. In this section, the entry wall stability with regard factor thatpillar affects mining, the narrow entry coal stability. pillar mining In this and section, non‐pillar the entry mining wall by filling stability artificial with materials. regard to pillar width to pillar width was theoretically investigated in three frequently used mining patterns, i.e., wide coal was theoretically investigated in three frequently used mining patterns, i.e., wide coal pillar mining, pillar 2.3.1.mining, Wide narrow Coal Pillar coal Mining pillar mining and non‐pillar mining by filling artificial materials. narrow coal pillar mining and non-pillar mining by filling artificial materials. In the wide coal pillar mining, disturbances from the current mining panel have little influence 2.3.1. Wide Coal Pillar Mining 2.3.1. Wideon the Coal next Pillar mining Mining panel. When the current mining panel is working, the tail entry or head entry Infor the the wide next coalmining pillar panel mining, can be disturbances tunneled at thefrom same the time,current as shownmining in panel Figure have 8. This little type influence of mining pattern is especially suitable for the condition where many panels must be mined onIn the the next wide mining coal pillar panel. mining, When the disturbances current mining from thepanel current is working, mining the panel tail haveentry little or head influence entry on simultaneously. thefor next the mining next mining panel. panel When can the be current tunneled mining at the panel same is working,time, as shown the tail in entry Figure or head8. This entry type for of the nextmining mining pattern panel canis especially be tunneled suitable at the samefor the time, condition as shown where in Figure many8. Thispanels type must of mining be mined pattern is especiallysimultaneously.Sustainability suitable 2017 for, 9, 2079 the condition where many panels must be mined simultaneously.9 of 19

γH

K2γH γH K1γH

r Panel I la il p al co Bw e id W Tail entry

Panel II

Head entry

Mining direction Figure 8. Plane layout of the wide coal pillar mining pattern. Figure 8. Plane layout of the wide coal pillar mining pattern. When the pillar is wide, the coal pillar can be divided into four zones, i.e., fractured zone (I and I’), plastic zone (II and II’), elastic zone (III and III’) and original stress zone (IV and IV’), along the pillar width direction, as shown in Figure 9. In fractured and plastic zones, the side abutment pressure can be expressed as [29]: 2 fx 1sin   =(Nem ), (1) 10 1sin 

where N0 is the self‐bearing capacity of the coal pillar; m is the mining height; f is the friction coefficient between roof rock and coal seam; x is the distance from the entry rib;  is the internal friction angle of the coal. The pressure at the elastic zone can be expressed as: 1  2 =1 K0  H , (2) ebx(  x0 )

where H and  are the thickness and unit weight of the overburden; x0 is the range sum of

fractured and plastic zones; K0 and b are constant coefficients reflecting the coal and rock strength. The pressure at the original stress zone can be written as:

 3 = H . (3)

In this kind of mining pattern, the mining influence of panel I has little influence on the mining of panel I, and there is no pressure superposition on the coal pillar ( d  0 ). Thus, the maximum

pressure on the coal pillar ( KHIII,  ) is only induced by one mining panel, and the pillar is relatively stable. This agrees with the numerical simulation and laboratory test results. The design width of the wide coal pillar in this mining pattern can be written as [30]:

Bxxww12 w L w, (4)

where B w is the design width of the pillar; xw1 and xw2 are the widths of the plastic deformation zone caused by the current and next working panels (panels I and II), respectively, and can be expressed using Equation (5); and L w is the width of the kernel elastic zone (about two times of the mining height).

Sustainability 2017, 9, 2079 9 of 19

When the pillar is wide, the coal pillar can be divided into four zones, i.e., fractured zone (I and I’), plastic zone (II and II’), elastic zone (III and III’) and original stress zone (IV and IV’), along the pillar width direction, as shown in Figure9. In fractured and plastic zones, the side abutment pressure can be expressed as [29]: 2 f x 1 + sin ϕ σ = N e m ( ), (1) 1 0 1 − sin ϕ where N0 is the self-bearing capacity of the coal pillar; m is the mining height; f is the friction coefficient between roof rock and coal seam; x is the distance from the entry rib; ϕ is the internal friction angle of the coal. The pressure at the elastic zone can be expressed as:

 1  σ2 = 1 + K0 γH, (2) eb(x−x0) where H and γ are the thickness and unit weight of the overburden; x0 is the range sum of fractured and plastic zones; K0 and b are constant coefficients reflecting the coal and rock strength. The pressure at the original stress zone can be written as:

σ3 = γH. (3)

In this kind of mining pattern, the mining influence of panel I has little influence on the mining of panel I, and there is no pressure superposition on the coal pillar (d ≥ 0). Thus, the maximum pressure on the coal pillar (KI,II γH) is only induced by one mining panel, and the pillar is relatively stable. This agrees with the numerical simulation and laboratory test results. The design width of the wide coal pillar in this mining pattern can be written as [30]:

Bw ≥ xw1 + xw2 + Lw, (4) Sustainability 2017, 9, 2079 10 of 19 where Bw is the design width of the pillar; xw1 and xw2 are the widths of the plastic deformation zone C caused by the current and next workingKH panels (panels I and II), respectively, and can be expressed II, I using Equation (5); and Lwmmis the width of thetg kernel elastic zone (aboutKHII,I  two times of the2 mining height). xtgtgw1,w 2 ln ln(1 )  , (5) 22tgC tg C  " Cc # "  λ # mλ KI,II γH + tgϕ m K γH x = ln tg + ln (1 + I,II tgϕ) + tg2 ϕ , (5) w1,w2 2tgϕ C 2tgϕ C tgϕ + σc where  is the side pressure coefficient; C is the internal friction angle of the coal;  c is the where λ is the side pressure coefficient; C is the internal friction angle of the coal; σc is the uniaxial uniaxial compressive strength of the coal pillar; KI ,II is the stress concentration factor induced by compressive strength of the coal pillar; KI,II is the stress concentration factor induced by coal coal face mining. face mining.

σ KIγH KIIγH

γH Gob of panel I Gob of panel II

I II III IV IV' III' II' I' x x0 d x0 L

FigureFigure 9. 9.Stress Stress distributions distributions inin a wide coal coal pillar pillar mining mining pattern. pattern.

Generally,Generally, stabilities stabilities of of the the entry entry in in aa widewide coal pillar mining mining pattern pattern can can be be well well guaranteed. guaranteed. However,However, oversize oversize pillars pillars usually usually causecause largelarge amountsamounts of of waste, waste, accounting accounting for for approximately approximately 10%~30% of the total coal production. In addition, each mining panel requires two entries that are mainly used for transportation and ventilation, which also increases construction cost and labor.

2.3.2. Narrow Coal Pillar Mining The narrow coal pillar mining pattern is also known as gob‐side entry driving [31–33]. The recovery ratio of this type of mining pattern is much larger than that of the wide coal pillar mining pattern, as shown in Figure 10.

γH

K2γH

K1γH γH

r la il Panel I p al o c BN w ro ar Tail entry N Panel II

Head entry

Mining direction Figure 10. Plane layout of the narrow coal pillar mining pattern.

However, when the pillar is narrow, the side abutment pressures induced by two adjacent mining panels couple with each other, which causes the integral pressure to be larger than that of a wide coal pillar. In this kind mining pattern, there is no original stress zone, as shown in Figure 11. The coupled pressure on the coal pillar can be written as:

Sustainability 2017, 9, 2079 10 of 19

C KH  II, I mm tg  KHII,I   2  xtgtgw1,w 2 ln ln(1 )  , (5) 22tgC tg C   c   tg

where  is the side pressure coefficient; C is the internal friction angle of the coal;  c is the

uniaxial compressive strength of the coal pillar; KI ,II is the stress concentration factor induced by coal face mining.

σ KIγH KIIγH

γH Gob of panel I Gob of panel II

I II III IV IV' III' II' I' x x0 d x0 L

Figure 9. Stress distributions in a wide coal pillar mining pattern. Sustainability 2017, 9, 2079 10 of 19 Generally, stabilities of the entry in a wide coal pillar mining pattern can be well guaranteed. However, oversize pillars usually cause large amounts of waste, accounting for approximately 10%~30%10%~30% of of the the total total coal coal production.production. In In addition, addition, each each mining mining panel panel requires requires two entries two entries that are that aremainly mainly used used for for transportation transportation and and ventilation, ventilation, which which also also increases increases construction construction cost and cost labor. and labor.

2.3.2.2.3.2. Narrow Narrow Coal Coal Pillar Pillar Mining Mining TheThe narrow narrow coal coal pillarpillar mining pattern pattern is isalso also known known as gob as‐side gob-side entry entrydriving driving [31–33]. [ 31The–33 ]. Therecovery recovery ratio ratio of this of this type type of mining of mining pattern pattern is much is larger much than larger that than of the that wide of thecoal wide pillar coal mining pillar miningpattern, pattern, as shown as shown in Figure in Figure 10. 10.

γH

K2γH

K1γH γH

r la il Panel I p al o c BN w ro ar Tail entry N Panel II

Head entry

Mining direction

FigureFigure 10. 10.Plane Plane layout layout ofof thethe narrow coal pillar pillar mining mining pattern. pattern.

However,However, when when the the pillar pillar is narrow, is narrow, the sidethe side abutment abutment pressures pressures induced induced by two by adjacenttwo adjacent mining panelsmining couple panels with couple each with other, each which other, causes which the causes integral the pressureintegral pressure to be larger to be than larger that than of athat wide of coala pillar.Sustainabilitywide In coal this pillar. kind2017, 9 mining, In2079 this kind pattern, mining there pattern, is no original there is stressno original zone, stress as shown zone, in as Figure shown 11 in. TheFigure11 coupled of 11. 19 pressureThe coupled on the pressure coal pillar on can the becoal written pillar can as: be written as: 11sin2 fx    1  2 f xm 1 + sin ϕ 30=1KHNe 0 ( ), (6) σ3 = 1+ K0 bL() x γH + N0e m ( ), (6) ebe(L−x) 11sin−sin ϕ L wherewhereL is the is width the width of the of coalthe coal pillar. pillar. As seen from Equation (6), as the width of the coal pillar decreases, the integral pressure As seen from Equation (6), as the width of the coal pillar decreases, the integral pressure increases increases accordingly, owing to the stress superposition effects. When the stress state of the coal accordingly, owing to the stress superposition effects. When the stress state of the coal meets the failure meets the failure criterion, the pillar will be damaged or crushed. The above theoretical model can criterion, the pillar will be damaged or crushed. The above theoretical model can well explain the well explain the numerical simulation and laboratory test results. numerical simulation and laboratory test results.

σ K'γH

KIγH KIIγH

γH

Gob of panel I Gob of panel II

III II' I' x x1 x1 L

FigureFigure 11. 11.Stress Stress distributions distributions in a narrow coal coal pillar pillar mining mining pattern. pattern.

TheThe size size of of the the pillar pillar is is difficult difficult to to bebe accuratelyaccurately determined for for a anarrow narrow coal coal pillar pillar mining mining pattern.pattern. If the If the width width of of the the pillar pillar is is irrational, irrational, the the mining-induced mining‐induced fractures fractures in in the the pillar pillar maymay leadlead toto air leakageair leakage and spontaneous and spontaneous combustion. combustion. In addition, In addition, the entrythe entry for thefor the next next mining mining panel panel is near is near the the peak peak position of the side abutment pressure in most cases, which may cause strong strata movements and large deformations. The width of the coal pillar for the NCPM pattern can be expressed by [34,35]:

BXXXN123, (7)

where X 1 is the width of the mining‐induced plastic area in the NCPM and can be given by  C  KH  m  tg  X 1  ln   , (8) 2tg C   c   tg 

X 2 is the effective anchorage length of the bolt (cable); X 3 is the additional safety length considering pillar stabilities, empirically approximately 30%~50% of （）XX12 ; and K is the stress concentration factor.

2.3.3. Non‐pillar Mining In development since the early 1950s, gob‐side entry retaining is still one of the most commonly used techniques in non‐pillar mining by filling artificial materials. In this technique, the former entry is artificially retained for the next mining panel by filling paste‐like backfill material, high‐water packing material and other artificial materials, as shown in Figure 12. The key technique of this mining pattern is that the narrow coal pillar is further replaced by the filling materials. However, the stress distribution on the artificial filing body is similar with that on a narrow coal pillar. The design width of the filling body can be written as [36]:

Sustainability 2017, 9, 2079 11 of 19 position of the side abutment pressure in most cases, which may cause strong strata movements and large deformations. The width of the coal pillar for the NCPM pattern can be expressed by [34,35]:

BN ≥ X1 + X2 + X3, (7) where X1 is the width of the mining-induced plastic area in the NCPM and can be given by

" C # mλ KγH + tgϕ X = ln , (8) 1 2tgϕ C tgϕ + σc

X2 is the effective anchorage length of the bolt (cable); X3 is the additional safety length considering pillar stabilities, empirically approximately 30%~50% of (X1 + X2); and K is the stress concentration factor.

2.3.3. Non-pillar Mining In development since the early 1950s, gob-side entry retaining is still one of the most commonly used techniques in non-pillar mining by filling artificial materials. In this technique, the former entry is artificially retained for the next mining panel by filling paste-like backfill material, high-water packing material and other artificial materials, as shown in Figure 12. The key technique of this mining pattern is thatSustainability the narrow 2017, 9 coal, 2079 pillar is further replaced by the filling materials. However, the stress distribution12 of 19 on the artificial filing body is similar with that on a narrow coal pillar. The design width of the filling body can be written as [36]: kFs BF  , kF (9) B ≥ s ,' (9) F σ0 where Fs is the roof pressure that acts on the filling body;  ' is the average compressive where F is the roof pressure that acts on the filling body; σ0 is the average compressive strength of the strengths of the filling body, and k is the safety factor, approximately 1.1~1.2. k filling body,The application and is the of safety this technique factor, approximately can reduce entry 1.1~1.2. drivage ratios and increase coal recovery rates.The However, application the of high this stress technique concentration can reduce on entrythe filling drivage body ratios and huge and increase material coal waste recovery problems rates. However,are always the a high concern. stress If concentration this type of mining on the pattern filling bodywere andused huge in a materialthick coal waste seam, problems the above are alwaysproblems a concern. would If be this more type serious. of mining Hence, pattern it is imperative were used to in find a thick a more coal effective seam, the and above economical problems wouldmethod be more to improve serious. entry Hence, stability it is and imperative mining sustainability. to find a more effective and economical method to improve entry stability and mining sustainability. γH

K2γH

K1γH γH

r la il Artificial filling body Panel I p B al F o c o Retained entry N Panel II

Head entry

Mining direction

FigureFigure 12. 12.Plane Plane layout layout of the non non-pillar‐pillar mining mining pattern. pattern.

3. Optimization3. Optimization of of the the Entry Entry Stability Stability in in Longwall Longwall Mining NumericalNumerical simulation simulation showed showed thatthat thethe coal pillar was was always always a astress stress-concentrated‐concentrated area area in in conventionalconventional longwall longwall mining. mining. Stress Stress transmissiontransmission induced induced by by panel panel mining mining was was the the main main reason reason accountingaccounting for for instability instability ofof the entry entry walls. walls. Laboratory Laboratory tests tests indicated indicated that a that wide a coal wide pillar coal was pillar wasfavorable favorable for for entry entry stabilities, stabilities, but but substantial substantial waste waste of ofcoal coal resources resources was was a major a major concern. concern. Although mining with narrow pillars can save more resources, the entry in this mining pattern Although mining with narrow pillars can save more resources, the entry in this mining pattern tended to be more unstable. Therefore, it can be concluded that coal pillar is one of the vulnerable entry surroundings where instability occurs. In the past few decades, non‐pillar mining by filling artificial materials has been tested in many mines. However, the filling body is also a stress‐concentrated area just as is the narrow coal pillar [37–39]. Additionally, excessive consumption of filling materials also sets limits on the popularity of this kind of mining pattern [40,41]. In longwall mining, roof caving is a complex dynamic process involving rock fracturing, disintegration and stress redistribution. Because movements of the entry and gob roof are always associated in the conventional mining patterns, the stress transmission from the gob roof is a main cause for stress concentration. Considering this, an innovative approach to improve entry stability by canceling the coal pillar and interrupting the stress transmission was proposed.

3.1. Directional Roof Fracturing (DRF) Technique Before mining, the entry and gob roof compose the entire plate structure. After mining, the gob roof gradually caves due to the roof weighting. The movement of the gob roof inevitably imposes a pulling force on the entry roof, thus causing deformations in the entry roof. If the pulling force were reduced or prevented, the stress condition of the entry would be alleviated. Considering this, a stress optimization method by fracturing the roof was proposed, as shown in Figure 13. Unlike the none pillar mining by filling artificial materials, the entry in this method is retained utilizing the bulking roof rocks. After the roof is fractured, the stress transferring from the gob roof is disrupted.

Sustainability 2017, 9, 2079 12 of 19 tended to be more unstable. Therefore, it can be concluded that coal pillar is one of the vulnerable entry surroundings where instability occurs. In the past few decades, non-pillar mining by filling artificial materials has been tested in many mines. However, the filling body is also a stress-concentrated area just as is the narrow coal pillar [37–39]. Additionally, excessive consumption of filling materials also sets limits on the popularity of this kind of mining pattern [40,41]. In longwall mining, roof caving is a complex dynamic process involving rock fracturing, disintegration and stress redistribution. Because movements of the entry and gob roof are always associated in the conventional mining patterns, the stress transmission from the gob roof is a main cause for stress concentration. Considering this, an innovative approach to improve entry stability by canceling the coal pillar and interrupting the stress transmission was proposed.

3.1. Directional Roof Fracturing (DRF) Technique Before mining, the entry and gob roof compose the entire plate structure. After mining, the gob roof gradually caves due to the roof weighting. The movement of the gob roof inevitably imposes a pulling force on the entry roof, thus causing deformations in the entry roof. If the pulling force were reduced or prevented, the stress condition of the entry would be alleviated. Considering this, a stress optimizationSustainability 2017 method, 9, 2079 by fracturing the roof was proposed, as shown in Figure 13. Unlike the13 of none 19 pillar mining by filling artificial materials, the entry in this method is retained utilizing the bulking In this way, the coal pillar is entirely removed, and the pressure on the entry roof is roof rocks. After the roof is fractured, the stress transferring from the gob roof is disrupted. In this correspondingly reduced. Additionally, the gob roof caves into loose gangues that bear the whole way, the coal pillar is entirely removed, and the pressure on the entry roof is correspondingly reduced. overburden. Because the bearing range of the loose gangue body is much larger than that of the coal Additionally, the gob roof caves into loose gangues that bear the whole overburden. Because the pillar or artificial filling body, the stress concentration would be alleviated after adopting the bearingtechnique. range of the loose gangue body is much larger than that of the coal pillar or artificial filling body, the stress concentration would be alleviated after adopting the technique.

γH

K γH Cross-section view 2 K1γH γH Gangues from the gob roof Panel I Roof fracturing line Retained entry Fracturing line Cables Gob Panel II

Retained entry Ganuges Panel II Head entry

Mining direction

FigureFigure 13. 13.Principle Principle of of the the mining mining pattern pattern usingusing the directional roof roof fracturing fracturing (DRF) (DRF) technique. technique.

Presently,Presently, blasting blasting is primarily is primarily used used to conduct to conduct roof fracturing roof fracturing in the field. in the To reducefield. To disturbances reduce ondisturbances the entry roof, on the a directional entry roof, roofa directional fracturing roof (DRF) fracturing technique (DRF) istechnique proposed. is proposed. In this technique, In this thetechnique, blasting energy the blasting is actively energy controlled is actively using controlled an energy using an binding energy tube. binding The tube. detonation The detonation generates high-speed,generates high-temperaturehigh‐speed, high‐temperature and high-pressure and high gases‐pressure that intensivelygases that actintensively in the predetermined act in the direction,predetermined as shown direction, in Figure as 14 shown. Practice in Figure indicated 14. Practice that the indicated DRF technique that the could DRF not technique only protect could the entrynot roof only from protect blasting the entry disturbances, roof from blasting but also disturbances, be favorable but for also rock be bulking. favorable for rock bulking. (a) (b)

Random fractures

Fracturing line Entry roof Entry roof

Gob roof Gob roof

Energy accumulated device

Figure 14. Active control of the crack propagation using the DRF technique. (a) Conventional blasting method. The blasting‐induced cracks propagate in random directions in which case the entry roof might be damaged. (b) Directional blasting method. The blasting‐induced cracks propagate in the predetermined direction in which case the entry roof can be well protected.

3.2. Stability Analysis of the Entry Using the Directional Roof Fracturing (DRF) Technique

3.2.1. Numerical Simulation (1) Modeling scheme Because the finite difference method is based on continuum mechanics, there are limitations in simulating roof caving problems. In contrast, the discrete element method (DEM) is a dynamic numerical method. The blocks or elements in the DEM modeling can shift, rotate or even separate from the main model. Therefore, the DEM is more applicable in simulating discontinuous and large deformation problems [42,43]. In this study, the DEM was used to explore entry stabilities under different mining conditions. The boundary conditions and mechanical parameters of the DEM model were same as those in the

Sustainability 2017, 9, 2079 13 of 19

In this way, the coal pillar is entirely removed, and the pressure on the entry roof is correspondingly reduced. Additionally, the gob roof caves into loose gangues that bear the whole overburden. Because the bearing range of the loose gangue body is much larger than that of the coal pillar or artificial filling body, the stress concentration would be alleviated after adopting the technique.

γH

K γH Cross-section view 2 K1γH γH Gangues from the gob roof Panel I Roof fracturing line Retained entry Fracturing line Cables Gob Panel II

Retained entry Ganuges Panel II Head entry

Mining direction

Figure 13. Principle of the mining pattern using the directional roof fracturing (DRF) technique.

Presently, blasting is primarily used to conduct roof fracturing in the field. To reduce disturbances on the entry roof, a directional roof fracturing (DRF) technique is proposed. In this technique, the blasting energy is actively controlled using an energy binding tube. The detonation generates high‐speed, high‐temperature and high‐pressure gases that intensively act in the Sustainabilitypredetermined2017, 9, 2079 direction, as shown in Figure 14. Practice indicated that the DRF technique could13 of 19 not only protect the entry roof from blasting disturbances, but also be favorable for rock bulking.

(a) (b)

Random fractures

Fracturing line Entry roof Entry roof

Gob roof Gob roof

Energy accumulated device

FigureFigure 14. 14.Active Active control control of the of crackthe crack propagation propagation using using the DRF the technique.DRF technique. (a) Conventional (a) Conventional blasting method.blasting The method. blasting-induced The blasting cracks‐induced propagate cracks propagate in random in directionsrandom directions in which in case which the case entry the roof mightentry be roof damaged. might (beb) Directionaldamaged. ( blastingb) Directional method. blasting The blasting-induced method. The blasting cracks‐induced propagate cracks in the predeterminedpropagate in directionthe predetermined in which direction case the entryin which roof case can the be entry well protected.roof can be well protected.

3.2.3.2. Stability Stability Analysis Analysis of of the the Entry Entry Using Using the the DirectionalDirectional Roof Fracturing Fracturing (DRF) (DRF) Technique Technique

3.2.1.3.2.1. Numerical Numerical Simulation Simulation (1)(1) Modeling Modeling scheme scheme BecauseBecause the the finite finite difference difference method is is based based on on continuum continuum mechanics, mechanics, there there are limitations are limitations in in simulatingsimulating roof roof caving caving problems. problems. In In contrast, contrast, the the discrete discrete element element method method (DEM) (DEM) is a is dynamic a dynamic numerical method. The blocks or elements in the DEM modeling can shift, rotate or even separate numerical method. The blocks or elements in the DEM modeling can shift, rotate or even separate from the main model. Therefore, the DEM is more applicable in simulating discontinuous and large from the main model. Therefore, the DEM is more applicable in simulating discontinuous and large deformation problems [42,43]. deformation problems [42,43]. In this study, the DEM was used to explore entry stabilities under different mining conditions. In this study, the DEM was used to explore entry stabilities under different mining conditions. The boundary conditions and mechanical parameters of the DEM model were same as those in the The boundary conditions and mechanical parameters of the DEM model were same as those in the preceding FDM model. Three modeling schemes were compared in the simulation, as presented in Figure 15. The narrow pillar mining pattern with a 5-m coal pillar, which is commonly used in many mines, was first simulated. Subsequently, stabilities of the entry in the non-pillar mining using the DRF technique were studied. To highlight the effects of the DRF technique, a non-pillar mining condition without using the DRF technique has also been simulated. It is generally known that roof deformation is an important factor describing entry stabilities. To quantify the stabilities of the entry in different mining conditions, monitoring points were respectively established at the middle of the entry roof. (2) Stability analysis of the entry in the simulation Figure 15 shows the structural morphologies of the entry and vertical displacement of the roof after mining. In the narrow coal pillar mining pattern, there was a long hanging roof structure above the entry wall (pillar), which easily caused stress concentrations and deformations. When the strength of the pillar cannot bear the overburden, an instability phenomenon would occur. As seen from Figure 15a, the entry wall has obviously deformed under the above heavy loads. The maximum sagging value of the entry roof was approximately 835 mm in the first scheme. Sustainability 2017, 9, 2079 14 of 19

preceding FDM model. Three modeling schemes were compared in the simulation, as presented in Figure 15. The narrow pillar mining pattern with a 5‐m coal pillar, which is commonly used in many mines, was first simulated. Subsequently, stabilities of the entry in the non‐pillar mining using the DRF technique were studied. To highlight the effects of the DRF technique, a non‐pillar mining condition without using the DRF technique has also been simulated. It is generally known that roof deformation is an important factor describing entry stabilities. To quantify the stabilities of the entry in different mining conditions, monitoring points were respectively established at the middle of the entry roof. (2) Stability analysis of the entry in the simulation Figure 15 shows the structural morphologies of the entry and vertical displacement of the roof after mining. In the narrow coal pillar mining pattern, there was a long hanging roof structure above the entry wall (pillar), which easily caused stress concentrations and deformations. When the strength of the pillar cannot bear the overburden, an instability phenomenon would occur. As seen Sustainabilityfrom Figure2017, 915a,, 2079 the entry wall has obviously deformed under the above heavy loads. The14 of 19 maximum sagging value of the entry roof was approximately 835 mm in the first scheme.

(a) (b) (c)

FigureFigure 15. 15.Structural Structural morphologies morphologies and displacementand displacement variation variation under differentunder different simulation simulation conditions. conditions. (a) The narrow coal pillar mining condition; (b) Non‐pillar mining using the DRF (a) The narrow coal pillar mining condition; (b) Non-pillar mining using the DRF technique; (c) technique; (c) Non‐pillar mining without using the DRF technique. Non-pillar mining without using the DRF technique.

ComparedCompared with with the the narrow narrow coal coal pillar pillar mining mining pattern,pattern, the roof subsidence decreased decreased about about 66% 66% to 283 mm after adopting the DRF technique. After fracturing, the motion state of the entry and gob to 283 mm after adopting the DRF technique. After fracturing, the motion state of the entry and gob roof became less associated. The gob roof caved into gangues along the fracturing line. The bulking roof became less associated. The gob roof caved into gangues along the fracturing line. The bulking gangues effectively supported the main roof, which helped decrease the amount of roof sagging. gangues effectively supported the main roof, which helped decrease the amount of roof sagging. Furthermore, if the roof was not fractured, the entry roof collapsed together with the gob roof, as Furthermore,shown in Figure if the 15c. roof The was amount not fractured, of roof sagging the entry at the roof monitoring collapsed point together was with approximately the gob roof, as1891 shown mm in in Figure the last 15 scheme.c. The From amount the simulation of roof sagging results, at especially the monitoring the second point and was last schemes, approximately we 1891can mm understand in the last that scheme. the roof From fracturing the simulation exerted a results,significant especially influence the on second the movement and last of schemes, the wesurrounding can understand rocks. that The the entry roof roof fracturing was not only exerted in a alow significant stress environment influence but on thealso movement deformed less of the surroundingafter taking rocks. the new The technique. entry roof In was a certain not only geological in a low condition, stress environment entry stabilities but also using deformed the DRF less aftertechnique taking was the newbetter technique. than that using In a the certain conventional geological mining condition, method. entry stabilities using the DRF technique was better than that using the conventional mining method. 3.2.2. Field Test 3.2.2. Field Test (1). Test Process (1) Test Process Roof fracturing height is one of the key parameters in the DRF technique. The fracturing height should be sufficient to fill the mining room with bulking gangues [44]. The fracturing height can be calculated by: h − h − h H = c s h , (10) k − 1 where hc is the mining height; hs is the amount of floor sagging; hh is the amount of bottom bulging; and k is the bulking coefficient of the gob roof. Based on the actual condition of the S1201 mining panel in the Ningtiaota coal mine, the bulking coefficient of the gob roof is approximately 1.39. The average amounts of bottom bulging and floor sagging were 310 mm and 360 mm, respectively. Considering the theoretical result and field condition, the fracturing height was finally determined to be 9.0 m. Test processes of the DRF technique were as follows (see Figure 16). Sustainability 2017, 9, 2079 15 of 19

Roof fracturing height is one of the key parameters in the DRF technique. The fracturing height should be sufficient to fill the mining room with bulking gangues [44]. The fracturing height can be calculated by: hhh H  csh, (10) k 1

where hc is the mining height; hs is the amount of floor sagging; hh is the amount of bottom bulging; and k is the bulking coefficient of the gob roof. Based on the actual condition of the S1201 mining panel in the Ningtiaota coal mine, the bulking coefficient of the gob roof is approximately 1.39. The average amounts of bottom bulging and floor sagging were 310 mm and 360 mm, respectively. Considering the theoretical result and Sustainability 2017, 9, 2079 15 of 19 field condition, the fracturing height was finally determined to be 9.0 m. Test processes of the DRF technique were as follows (see Figure 16). (i) (i) DrillDrill fracturing fracturing holes. holes. TheThe fracturing holes holes were were marked marked on on the the roof roof in in advance. advance. The The hole hole diameterdiameter and and spacing spacing were 46 46 mm mm and and 600 600 mm, mm, respectively. respectively. To reduce To reduce disturbances disturbances from the from thegob gob roof roof caving, caving, the the fracturing fracturing angle angle must must be beconsistent, consistent, tilting tilting ten ten degrees degrees towards towards the the mined-outmined‐out area. area. (ii)(ii)Install Install energy energy binding binding tubes.tubes. First, the the energy energy binding binding tubes tubes were were charged charged with with emulsified emulsified blastingblasting powders powders outside outside the the fracturing fracturing holes. holes. Subsequently, Subsequently, the tubesthe tubes together together with thewith powders the werepowders put into were the put holes. into Notethe holes. that theNote energy-accumulated that the energy‐accumulated direction direction must be same,must be in same, which in case a continuouswhich case a fracturing continuous line fracturing could be line formed. could be formed. (iii) Perform roof fracturing. After blasting, the cracks formed a fracturing line on the surface of the (iii) Perform roof fracturing. After blasting, the cracks formed a fracturing line on the surface of the entry roof. As the mining panel advanced, the gob roof gradually caved along the fracturing entry roof. As the mining panel advanced, the gob roof gradually caved along the fracturing line, line, and the caved gangues formed a special entry wall (gangue rib). and the caved gangues formed a special entry wall (gangue rib).

FigureFigure 16. 16.Test Test processes processes of theof the DRF DRF technique. technique. (a) Drilling(a) Drilling fracturing fracturing holes; holes; (b) Charging(b) Charging with with energy bindingenergy tubes; binding (c) tubes; Roof fracturing (c) Roof fracturing effects. effects.

(2) Stability Analysis of the Entry in the Field (2) Stability Analysis of the Entry in the Field To explore the influence of the DRF technique in the field, a KJ216 pressure monitoring apparatusTo explore was the installed influence on of the the face DRF‐end technique support in beside the field, the aentry, KJ216 as pressure shown monitoringin Figure 17. apparatus This wasapparatus installed could on the continuously face-end support and automatically beside the entry, record as shown the roof in Figurepressure 17 .during This apparatus the mining could continuouslyprocess. The and monitoring automatically distances record of the the non roof‐fracturing pressure and during fracturing the mining condition process. were both The monitoring150 m. distancesFigure of the 18 non-fracturing shows the variation and fracturingof the roof conditionpressure before were and both after 150 adopting m. the DRF technique. It canFigure be seen18 shows that the the maximum variation roof of the pressure roof pressure was 19.5 before MPa after and afterroof fracturing, adopting the8.7 DRFMPa smaller technique. It canthan be that seen in thatthe non the‐fracturing maximum condition. roof pressure The average was 19.5 roof MPa pressure after roofdecreased fracturing, by appropriately 8.7 MPa smaller 60% thanto that7.1 MPa in the after non-fracturing adopting the condition. new technique. The average Field monitoring roof pressure indicated decreased that bythe appropriatelyDRF technique 60% to 7.1could MPa effectively after adopting optimize the newthe technique.stress condition Field monitoringof the entry. indicated This would that thealso DRF be techniquefavorable couldto effectivelystabilities optimize and safeties the of stress the entry condition in follow of the‐up entry.maintenance. This would also be favorable to stabilities and safetiesSustainability of the 2017 entry, 9, 2079 in follow-up maintenance. 16 of 19

150 m 150 m Before fracturing After fracturing Face-end support

S1201 headentry

S1201 mining panel Gob

(a) (b)

FigureFigure 17. 17.Sketch Sketch of of the the field field monitoring monitoring andand apparatus. ( (aa)) Monitoring Monitoring position position and and scope; scope;( b) KJ216 monitoring(b) KJ216 apparatusmonitoring for apparatus roof pressure. for roof pressure.

40 40 Before fracturing After fracturing Max : 28.2 MPa 30 1 30

Max2: 19.5 MPa 20 20

10 10 Roof pressure (MPa) Roof pressure (MPa)

0 0 0 30 60 90 120 150 180 210 240 270 Mining distance/m Figure 18. Variation of the roof pressure before and after adopting the DRF technique.

Figure 19 shows the final entry retaining photos after adopting the DRF technique. The gob roof collapsed into gangues along the fracturing line. The caved gangues replaced the coal pillar or artificial filling body and became another entry rib. During caving and compaction of the gangues, gangue prevention structures and metal nets were used to prevent the gangues from extending out into the entry, as illustrated in Figure 19a. Once the gangues and main roof had been stabilized, air leakage prevention materials were sprayed on the surface of the gangue rib, as shown in Figure 19b. The mean sagging amount of the entry roof was approximately 258 mm. The cross‐section of the retained entry could fully meet requirements of the next mining panel. The field test revealed that the DRF technique could effectively improve stabilities of the entry.

Figure 19. Entry retaining effects after adopting the new technique. (a) Roof caving effects; (b) Entry retaining effects.

Sustainability 2017, 9, 2079 16 of 19 Sustainability 2017, 9, 2079 16 of 19 150 m 150 m Before150 fracturing m After150 fracturing m Face-end support Before fracturing After fracturing Face-end support S1201 headentry S1201 headentry

S1201 mining panel Gob S1201 mining panel Gob

(a) (b) (a) (b) SustainabilityFigure2017 17., 9Sketch, 2079 of the field monitoring and apparatus. (a) Monitoring position and scope; 16 of 19 (Figureb) KJ216 17. monitoring Sketch of the apparatus field monitoring for roof pressure. and apparatus. (a) Monitoring position and scope; (b) KJ216 monitoring apparatus for roof pressure. 40 40 40 Before fracturing After fracturing 40 Before fracturing After fracturing Max1: 28.2 MPa 30 Max : 28.2 MPa 30 30 1 30 Max2: 19.5 MPa 20 20 Max2: 19.5 MPa 20 20

10 10 Roof pressure (MPa)

Roof pressure (MPa) 10 10 Roof pressure (MPa) Roof pressure (MPa) 0 0 00 30 60 90 120 150 180 210 240 270 0 0 30 60 90Mining 120 distance/m 150 180 210 240 270 Mining distance/m FigureFigure 18. 18. VariationVariation of the roof roof pressure pressure before before and and after after adopting adopting the the DRF DRF technique. technique. Figure 18. Variation of the roof pressure before and after adopting the DRF technique. Figure 19 shows the final entry retaining photos after adopting the DRF technique. The gob roofFigure collapsedFigure 19 19 shows intoshows gangues the the final final along entry entry the retaining retainingfracturing photos photos line. The after after caved adopting adopting gangues the the DRFreplaced DRF technique. technique. the coal The pillarThe gob gob or roof collapsedartificialroof collapsed intofilling gangues intobody gangues and along became thealong fracturing another the fracturing entry line. rib. The line. During caved The caved gangues caving gangues and replaced compaction replaced the coal the of pillarthecoal gangues, pillar or artificial or fillinggangueartificial body prevention filling and body became structures and another became and entry anothermetal rib. nets entry During were rib. used During caving to prevent caving and compaction theand gangues compaction of from the of extending gangues, the gangues, gangueout preventionintogangue the preventionentry, structures as illustrated structures and metal in andFigure nets metal were19a. nets Once used were the to used preventgangues to prevent theand gangues main the ganguesroof from had extendingfrombeen extendingstabilized, out into outair the entry,leakageinto asthe illustrated preventionentry, as illustrated inmaterials Figure in were19 Figurea. Oncesprayed 19a. the Onceon gangues the the surface gangues and of main the and gangue roof main had roof rib, been ashad shown stabilized,been stabilized,in Figure air leakage19b. air preventionTheleakage mean prevention materials sagging amountmaterials were sprayed of were the on entrysprayed the roof surface on was the of approximatelysurface the gangue of the rib,gangue 258 as mm. shown rib, The as in showncross Figure‐section in 19Figureb. of The 19b.the mean saggingretainedThe mean amount entry sagging could of the amount fully entry meet of roof the requirements was entry approximately roof wasof the approximately next 258 mining mm. The panel.258 cross-section mm. The The field cross test of the‐ sectionrevealed retained of that the entry couldtheretained DRF fully techniqueentry meet could requirements could fully effectively meet of requirements the improve next mining stabilities of the panel. next of The themining fieldentry. panel. test revealed The field that test the revealed DRF technique that couldthe DRF effectively technique improve could effectively stabilities ofimprove the entry. stabilities of the entry.

Figure 19. Entry retaining effects after adopting the new technique. (a) Roof caving effects; (b) Entry FigureretainingFigure 19. 19. effects. EntryEntry retaining retaining effects effects after after adopting adopting the new the technique. new technique. (a) Roof caving (a) Roof effects; caving (b) Entry effects; (bretaining) Entry retaining effects. effects.

4. Conclusions In China, coal provides most of the energy for the development of economy and industrialization. Mining depth is gradually increasing due to limited coal resources in shallow depth. Exploiting the valuable coal resources efficiently and safely is meaningful to the sustainable development of resources and energy, as well as the economy and society. In this paper, the stability of the entry is selected as the research target. The main conclusions in this paper are listed as follows.

(1) Simulations of different mining patterns showed that mining activities could cause stress concentrations on the entry wall. When the width of the entry wall was more than 30 m, there were two stress peaks distributed on the sides. As the width decreased, the stress peaks moved closer to each other and gradually merged into one. During this process, the vertical stresses caused by two mining panels, coupled with each other, and the integrated stress gradually increased. However, when the width was less than 10 m, the coupling stress decreased unexpectedly. Sustainability 2017, 9, 2079 17 of 19

(2) Entry wall stabilities were simply explored using representative coal units at the laboratory. True triaxial tests showed that, as the confining stress decreased, the critical buckling load decreased, and the degree of damage became more serious. The testing results implied that a wide pillar was more stable than a narrow one. Nevertheless, the serious waste of resources was a considerable concern for mining with wide coal pillars. Hence, mining with coal pillars always involves a tradeoff between economy and safety. (3) Considering the existing problems of longwall mining with coal pillars, an innovative approach to improve entry stabilities by optimizing the stress distributions was introduced. In this approach, the geometry of the entry and gob roofs is actively controlled by roof fracturing along with support, thus reducing stress in the retained entry. Additionally, bulking characteristics of the caving gangues at the gob area are efficiently used, and no extra man-made filling material is therefore needed, which makes this approach more flexible and cost effective compared with other previous approaches. Numerical modeling showed that the DRF technique could prevent stress transfer from the gob roof to entry roof, thus causing the entry to be more stable. The deformation of the entry roof decreased by approximately 66% compared with that using the conventional mining method. Field tests indicated that the average roof pressure decreased by approximately 60% after adopting the new technique. Together, these results suggested that the DRF technique could not only avoid resource waste but also be favorable to improving entry stabilities.

Acknowledgments: This work is supported by the National Natural Science Foundation of China (No. 51704298, No. 51674265) and the State Key Research Development Program of China (No. 2016YFC0600900), which are gratefully acknowledged. Author Contributions: Manchao He and Yubing Gao proposed the new technique and conceived the experiments; Yubing Gao and Xingyu Zhang performed the numerical simulation and field tests; Yubing Gao and Dongqiao Liu performed the laboratory experiments; Yubing Gao analyzed the data and wrote the paper. Conflicts of Interest: The authors declare no conflict of interest.

References

1. Milici, R.C.; Flores, R.M.; Stricker, G.D. Coal resources, reserves and peak coal production in the United States. Int. J. Coal Geol. 2013, 113, 109–115. [CrossRef] 2. Lechner, A.M.; Kassulke, O.; Unger, C. Spatial assessment of open cut coal mining progressive rehabilitation to support the monitoring of rehabilitation liabilities. Resour. Policy 2016, 50, 234–243. [CrossRef] 3. Shabanimashcool, M.; Li, C.C. A numerical study of stress changes in barrier pillars and a border area in a longwall coal mine. Int. J. Coal Geol. 2013, 106, 39–47. [CrossRef] 4. Suchowerska, A.M.; Merifield, R.S.; Carter, J.P. Vertical stress changes in multi-seam mining under supercritical longwall panels. Int. J. Rock Mech. Min. Sci. 2013, 61, 306–320. [CrossRef] 5. Basarir, H.; Ferid Oge, I.; Aydin, O. Prediction of the stresses around main and tail gates during top coal caving by 3D numerical analysis. Int. J. Rock Mech. Min. Sci. 2015, 76, 88–97. [CrossRef] 6. Su, D.W. Effects of longwall-induced stress and deformation on the stability and mechanical integrity of shale gas wells drilled through a longwall abutment pillar. Int. J. Min. Sci. Technol. 2017, 27, 115–120. [CrossRef] 7. Gao, Y.; Liu, S.; Lyu, B.; Li, K. Mechanism study of floor water inrush around mining field based on micro-crack extension. J. Min. Saf. Eng. 2016, 33, 624–629. 8. Cheng, G.; Ma, T.; Tang, C.; Liu, H.; Wang, S. A zoning model for coal mining—Induced strata movement based on microseismic monitoring. Int. J. Rock Mech. Min. Sci. 2017, 94, 123–138. [CrossRef] 9. Tu, S.; Wang, F.; Dou, F.; Yuan, Y.; Lu, Y. Fully mechanized top-coal caving: Underground stress at gate ways under barrier pillars of an upper coal seam. J. China Univ. Min. Technol. 2010, 39, 1–5. 10. Kang, H.; Wu, Y.; Gao, F. Deformation characteristics and reinforcement technology for entry subjected to mining-induced stresses. J. Rock Mech. Geotech. Eng. 2011, 3, 207–219. [CrossRef] 11. Wei, G. Study on the width of the non-elastic zone in inclined coal pillar for strip mining. Int. J. Rock Mech. Min. Sci. 2014, 72, 304–310. [CrossRef] Sustainability 2017, 9, 2079 18 of 19

12. Wang, S.; Hao, S.; Chen, Y.; Bai, J.; Wang, X.; Xu, Y. Numerical investigation of coal pillar failure under simultaneous static and dynamic loading. Int. J. Rock Mech. Min. Sci. 2016, 84, 59–68. [CrossRef] 13. Reed, G.; Mctyer, K.; Frith, R. An assessment of coal pillar system stability criteria based on a mechanistic evaluation of the interaction between coal pillars and the overburden. Int. J. Min. Sci. Technol. 2017, 27, 9–15. [CrossRef] 14. Wattimena, R.K.; Kramadibrata, S.; Sidi, I.D.; Azizi, M.A. Developing coal pillar stability chart using logistic regression. Int. J. Rock Mech. Min. Sci. 2013, 58, 55–60. [CrossRef] 15. Das, A.J.; Mandal, P.K.; Ghosh, C.N.; Sinha, A. Extraction of locked-up coal by strengthening of rib pillars with FRP-A comparative study through numerical modelling. Int. J. Min. Sci. Technol. 2017, 27, 261–267. [CrossRef] 16. Yang, J.; Yu, S.; Tao, Z.; Sun, X.; Wang, D. On the deformation and failure characteristics of the tertiary soft rock entry and coupling control measures. J. Min. Saf. Eng. 2014, 31, 373–378. 17. He, M.; Gao, Y.; Yang, J.; Guo, Z.; Wang, E.; Wang, Y. An energy-gathered roof cutting technique in no-pillar mining and its impact on stress variation in surrounding rocks. Chin. J. Rock Mech. Eng. 2017, 36, 1314–1325. 18. Zhang, N.; Yuan, L.; Han, C.; Xue, J.; Kan, J. Stability and deformation of surrounding rock in pillarless gob-side entry retaining. Saf. Sci. 2012, 50, 593–599. [CrossRef] 19. Islavath, S.R.; Deb, D.; Kumar, H. Numerical analysis of a longwall mining cycle and development of a composite longwall index. Int. J. Rock Mech. Min. Sci. 2016, 89, 43–54. [CrossRef] 20. Shabanimashcool, M.; Li, C.C. Numerical modelling of longwall mining and stability analysis of the gates in a coal mine. Int. J. Rock Mech. Min. Sci. 2012, 51, 24–34. [CrossRef] 21. Cheng, Y.; Wang, J.; Xie, G.; Wei, W. Three-dimensional analysis of coal barrier pillars in tailgate area adjacent to the fully mechanized top caving mining face. Int. J. Rock Mech. Min. Sci. 2010, 47, 1372–1383. [CrossRef] 22. Cai, M.; He, M.; Liu, D. Rock Mechanics and Engineering; Science Press: Beijing, China, 2013; pp. 32–38. 23. Wang, G.; Wu, M.; Wang, R.; Xu, H.; Song, X. Height of the mining-induced fractured zone above a coal face. Eng. Geol. 2017, 216, 140–152. [CrossRef] 24. Jing, Y.; Armstrong, R.T.; Mostaghimi, P. Rough-walled discrete fracture network modelling for coal characterisation. Fuel 2017, 191, 442–453. [CrossRef] 25. Su, F.; Itakura, K.; Deguchi, G.; Ohga, K. Monitoring of coal fracturing in underground coal gasification by acoustic emission techniques. Appl. Energy 2017, 189, 142–156. [CrossRef] 26. Huang, Z.; Dai, X.; Dong, L. Buckling failures of reserved thin pillars under the combined action of in-plane and lateral hydrostatic compressive forces. Comput. Geotech. 2017, 87, 128–138. [CrossRef] 27. Chen, S.; Guo, W.; Cheng, G.; Zhao, T. Research on creep supporting effect of deep strip pillar. J. Min. Saf. Eng. 2012, 29, 48–53. 28. Song, Y.; Yang, X. Evolution characteristics of deformation and energy fields during coal pillar instability. J. Min. Saf. Eng. 2013, 30, 822–827. 29. Gu, S.; Jiang, B.; Wang, C.; Li, N. Simulation research on overburden failure and lead abutment pressure distribution of fully-mechanized sublevel caving face. Min. Res. Dev. 2013, 33, 11–14. 30. Zhang, W.; Kong, X.; Kang, T.; Zhao, G. Research on reasonable size of large coal pillar between panels at mining face with soft surrounding rock. Min. Res. Dev. 2013, 33, 14–17. 31. Zhang, Y.; Wan, Z.; Li, F.; Zhou, C.; Zhang, B.; Guo, F.; Zhu, C. Stability of coal pillar in gob-side entry driving under unstable overlying strata and its coupling support control technique. Int. J. Min. Sci. Technol. 2013, 23, 193–199. [CrossRef] 32. Song, Z.; Cui, Z.; Xia, H.; Tang, J.; Wen, Z. The fundamental theoretical and engineering research on the green safe no coal pillar mining model by mainly using coal gangue backfill. J. China Coal Soc. 2010, 35, 705–710. 33. Hou, C.; Li, X. Stability principle of big and small structures of rock surrounding entry driven along goaf in fully mechanized top coal caving face. J. China Coal Soc. 2001, 26, 1–7. 34. Zhang, K.; Jiang, Y.; Zhang, Z.; Zhang, Y.; Pang, X.; Zeng, X. Determining the reasonable width of narrow pillar of entry in gob entry driving in the large pillar. J. Min. Saf. Eng. 2014, 31, 255–262. 35. Wang, D.; Li, S.; Wang, Q.; Li, W.; Wang, F.; Wang, H.; Peng, P.; Ruan, G. Experimental study of reasonable coal pillar width in fully mechanized top coal caving face of deep thick coal seam. Chin. J. Rock Mech. Eng. 2014, 33, 539–548. 36. Huang, Y.; Sun, H. Design of gob-side packing parameters for gateways maintained along the goaf. J. China Coal Soc. 1997, 22, 127–131. Sustainability 2017, 9, 2079 19 of 19

37. Zhang, J.; Li, B.; Zhou, N.; Zhang, Q. Application of solid backfilling to reduce hard-roof caving and longwall coal face burst potential. Int. J. Rock Mech. Min. Sci. 2016, 88, 197–205. [CrossRef] 38. Zhang, Z.; Bai, J.; Chen, Y.; Yan, S. An innovative approach for gob-side entry retaining in highly gassy fully-mechanized longwall top-coal caving. Int. J. Rock Mech. Min. Sci. 2015, 80, 1–11. [CrossRef] 39. Zhang, G.; He, M.; Yu, X.; Huang, Z. Research on the technique of no-pillar mining with gob-side entry formed by advanced roof caving in the protective seam in Baijiao coal mine. J. Min. Saf. Eng. 2011, 28, 511–516. 40. Bu, T.; Feng, G.; Jia, K. Gateway side backfilling support technology of goaf side gateway in fully mechanized high cutting longwall mining face. Coal Sci. Technol. 2010, 38, 41–44. 41. Ma, L.; Zhang, D.; Wang, H.; Li, Y. Mining technique with preset packing body in entry for thick coal seam without coal pillars. Chin. J. Rock Mech. Eng. 2010, 29, 674–680. 42. Leclerc, W.; Haddad, H.; Guessasma, M. On the suitability of a Discrete Element Method to simulate cracks initiation and propagation in heterogeneous media. Int. J. Solids Struct. 2017, 108, 98–114. [CrossRef] 43. Gui, Y.L.; Zhao, Z.Y.; Kodikara, J.; Bui, H.H.; Yang, S.Q. Numerical modelling of laboratory soil desiccation cracking using UDEC with a mix-mode cohesive fracture model. Eng. Geol. 2016, 202, 14–23. [CrossRef] 44. Gao, Y.; Guo, Z.; Yang, J.; Wang, J.; Wang, Y. Steady analysis of gob-side entry retaining formed by roof fracturing and control techniques by optimizing mine pressure. J. China Coal Soc. 2017, 42, 1672–1681.

© 2017 by the authors. Licensee MDPI, Basel, Switzerland. This article is an open access article distributed under the terms and conditions of the Creative Commons Attribution (CC BY) license (http://creativecommons.org/licenses/by/4.0/).