Abstract
Aiming at the problem of unclear crack propagation law of overburden rock in the process of small coal pillar mining in inclined extra-thick coal seam, this paper takes the 11,002 working face of Tengda Coal Mine in Heze, Xinjiang as the background. Through physical similarity simulation and 3DEC numerical simulation, the mining-related laws of small coal pillar working face in inclined extra-thick coal seam are clarified, and the height of caving zone and fracture zone are determined by combining multi-method weights. The results show that: (1) when mining in strike (dip), the fractal dimension of overburden fracture decreases rapidly in the range of 3–25 m (3–22 m) from the roof, the height of caving zone is 25 m, the height of fracture zone is 66.5 m (81.7 m), and the stress concentration coefficient is 1.17 in both cases; (2) The caving form of the whole overlying strata in the goaf is asymmetrically distributed, and a stable area is formed on the upper side of the small coal pillar, and the overlying strata do not collapse; (3) The subsidence and fracture rate of overlying strata show a vertical gradient decreasing law; (4) Using the weighted average method based on weight, the height of the caving zone is determined to be 24.98 m, and the height of the fracture zone is 81.67 m. The research results provide a theoretical basis for mining overburden control and gas control in small coal pillar working face of inclined coal seam.
Keywords: Inclined thick coal seam, Small coal pillar working face, Overburden rock fissures, Numerical simulation, Fracture evolution
Subject terms: Energy science and technology, Engineering, Natural hazards, Solid Earth sciences
Introduction
With the gradual depletion of coal resources in eastern China, the coal strategic base has gradually moved westward. The main occurrence conditions of coal seams in Xinjiang, the main coal base in the west, have the characteristics of large dip angle and extra thickness. Due to the difference and space–time of mining-induced overburden cracks in small coal pillar working face of inclined extra-thick coal seam, the layout of gas extraction system is mismatched with the actual location of mining-induced pressure relief gas high permeability area.
With the gradual depletion of coal resources in eastern China, the coal strategic base has gradually moved westward. The main occurrence conditions of coal seams in Xinjiang, the main coal base in the west, have the characteristics of large dip angle and extra thickness. Due to the difference and space–time of mining-induced overburden cracks in small coal pillar working face of inclined extra-thick coal seam, the layout of gas extraction system is mismatched with the actual location of mining-induced pressure relief gas high permeability area.
In terms of theoretical analysis, the fracture evolution law of overlying strata in coal seam mining has been widely concerned by scholars at home and abroad. Through the combination of physical simulation and numerical simulation, it has accumulated fruitful results1–3. These theories are mainly designed for the specific conditions of horizontal and near-horizontal coal seams, and their application scope and applicable scenarios have limitations. These theories have played a fundamental supporting role in promoting the construction and development of the theoretical system of strata control. In recent years, the academic community has made significant progress in exploring the complex geomechanical challenges encountered in the mining process of extra-thick coal seams. These advances have focused on many key fields such as the evolution law of overburden structure, the control strategy of surrounding rock stability and the technology of disaster prevention and control, which have significantly promoted the theoretical system and practical application in this field. Lv et al. proposed a two-stage mining method of “middle-level large-height mining and lower-level top-coal caving”. The dynamic influence mechanism of different parameters on top-coal migration and pressure distribution was analyzed by numerical simulation system. The results show that mining stress can be used to effectively promote top-coal crushing and smooth caving, so as to achieve efficient mining4. Aiming at the problem of surrounding rock control of roadway in extra-thick coal seam, through field measurement and simulation analysis, some scholars put forward the collaborative control technology with “optimizing coal pillar width” and “roof thick anchorage and high pre-tightening force support” as the core, which effectively guarantees the stability of roadway and improves the tunneling efficiency5,6. Aiming at the problem of surrounding rock control in deep and extra-thick coal seams, Yang et al. established a mechanical model of overlying strata before fracture, revealed the synergistic mechanism of mining face width controlling the fracture evolution of overlying strata and the deformation of soft rock roadway, and proposed a synergistic support technology with grouting anchor cable and concrete-filled steel tube support as the core, which improved the stability and control effect of surrounding rock under complex mining conditions7,8. Gao et al. focused on the mechanism of strong ground pressure induced by multi-layer hard roof fracture in fully-mechanized caving mining of ultra-thick coal seams. Through the secondary development of ABAQUS software, the process of rock caving was simulated, and the bearing stress distribution, fracture morphology and impact characteristics of hard roofs in different layers were revealed. The difference provides theoretical support for safe and efficient mining of extra-thick coal seams under hard roof conditions9. Hu et al. explored the evolution law of overburden fracture in variable mining height working face of deep buried thick coal seam by using physical similarity simulation experiment, and found that the fracture height showed a trend of “latent-sudden increase-gradual increase-stability”10. Aiming at the problem of gas outburst disaster prevention and control in deep coal seam group, Wang et al. constructed a three-dimensional gas extraction and utilization system, which can effectively reduce the risk of strong outburst coal seam and realize the efficient extraction and utilization of pressure relief gas. It provides a new way for gas control and resource development in deep coal seam in China11. Chen et al. used COMSOL multi-physical field coupling simulation to propose a large-diameter directional drilling treatment technology for high-concentration gas in fractured zones to achieve synergistic optimization of extraction efficiency improvement and cost control12. Zhu et al. aimed at the problem of strong mine pressure and roadway instability induced by tectonic stress in extra-thick coal seam mining. Through nonlinear dynamic mechanical model and physical simulation experiments, the evolution mechanism by which tectonic stress controls the scale of “plate and shell” structure of overlying strata and the formation of triangular support structure of overlying strata under mining action was revealed. The pre-splitting blasting technology and targeted support scheme proposed in the study have realized the effective control of strong mine pressure and roadway deformation, and provided theoretical basis and engineering countermeasures for safe mining of extra-thick coal seams13,14. Du et al.15 focused on the problem of resource recovery of wide coal pillars left over from layered mining of II1 coal seam in Zhao Gu second Coal Mine. Through numerical simulation and field industrial test, the research team revealed the mechanism of stress arch bearing structure above the section coal pillar during collaborative mining, deduced the working resistance formula of hydraulic support and determined reasonable parameters, which provided engineering reference for the collaborative mining of segmented coal pillars and lower coal seams. Through physical similarity simulation and numerical calculation, Lv et al. revealed the evolution characteristics of overburden fractures in paste composite filling mining, and found that fracture development was affected by factors such as mining height, filling rate and buried depth, which provided theoretical support for water-preserved mining16.
At present, domestic and foreign scholars have formed a systematic research system covering theory, design, support and monitoring around the technology of small coal pillar working face. Zhang et al. used numerical simulation method to study the width of small coal pillars in gob-side entry driving in shallow coal seams, determined that the reasonable width of coal pillars was 7.0 m, and proposed a sub-regional asymmetric combined support technology. Field application shows that this technology can significantly control the deformation of surrounding rock. Through theoretical calculation and measured comparison17. Feng et al. analyzed the dynamic evolution law of overburden structure in small coal pillar working face, determined that the maximum working resistance of support was 9153.48 kN, and verified that the resistance value could effectively guarantee the safe production of working face18. Based on the theory of stress field distribution, Chen et al. proposed a design method for the width of small coal pillars in gob-side entry driving. In the application of Zaoquan Coal Mine, 6.0 m was determined as the optimal width, and a complete area was observed in the coal pillar, which confirmed the reliability of the method19. Lin et al. used the split Hopkinson pressure bar system to carry out dynamic compression tests, and analyzed the law of energy input, dissipation and release under different strain rates. It was found that coal deformation showed significant strain rate dependence, and the dynamic strength and energy consumption parameters increased linearly with the strain rate, which provided theoretical support for the prevention and control of coal-rock dynamic disasters and efficient mining of resources20. Cao et al. used slurry grouting technology to repair mining fissures. Through systematic research, the diffusion law of grouting slurry under different working conditions is analyzed. The results show that under different grouting speeds, when the slurry diffuses to a certain distance, the pressure drops to the same level. It has important theoretical significance and engineering application value for grouting sealing effect21. Zhao et al. studied the stress distribution of the remaining coal pillars in the upper layer of the working face. By combining theoretical analysis, numerical simulation and field measurement, the stress variation law of the 8707 working face of 8# coal seam passing through the remaining coal pillars of 7# coal seam was obtained22.
The migration characteristics and path orientation of mining pressure relief gas are largely restricted by the spatial and temporal distribution characteristics of overburden fracture field. The research shows that the overlying strata in the process of mining failure shows the evolution law from bottom to top, from order to disorder, and this kind of overburden fracture field has significant nonlinear characteristics in essence23,24. It is worth noting that the superposition effect of periodic mine pressure appearance and repeated disturbance factors in actual mining operations further aggravates the complexity of the evolution process. Xie’s team revealed that the fracture network of mining overburden has unique self-similar fractal characteristics through systematic research. The fractal dimension theory was innovatively introduced into the field of rock mechanics research, and the mathematical quantitative characterization of random distribution and morphological irregularity of fractures was realized for the first time25. This breakthrough has promoted the follow-up research. Many scholars have used fractal theory to carry out quantitative analysis of fracture systems under different mining conditions, and achieved fruitful research results26.
Design of fracture evolution law of mining overburden rock in small coal pillar working face of inclined thick coal seam
General situation of test working face
The design strike length of the simulation experiment working face is 1320 m, the design tendency length is 176 m, the average coal thickness is 5 m, and the coal seam dip angle is 3°. Through laboratory analysis, the original gas content is 9.21 m3/t, the residual gas content is 7.18 m3/t, the gas emission is large, and the maximum can reach 74.35 m3/min. According to the coal seam occurrence, gas occurrence, working face conditions, etc., the gas emission amount of the working face can be predicted, and the gas extraction method of the mining fracture zone can be selected step by step, The test mine selects the strike high drainage roadway and the directional long borehole to synergistically extract and control gas.
Experimental scheme
According to the mining conditions of 11,002 small coal pillar fully mechanized caving face in Tengda Coal Mine, the experiment was carried out according to the similarity conditions and similarity criteria of physical simulation. In the physical similarity simulation experiment, the experimental model retains protective coal pillars on the left and right (100 m along the tendency, 10 m along the strike)”. According to the experimental similarity ratio of 1:200, when the strike mining is carried out, the mining height is set to 4.5 m, and the total advancing distance is130 m. When inclined mining, the mining height is set to 4.5 m, and the total advancing distance is 150 m, as shown in Fig. 1.
Fig. 1.
Two-dimensional physical similarity model.
Evolution law of mining overburden rock fracture
By analyzing the influence of mining on the overlying strata, it can be found that the continuous development of separation cracks and fracture cracks has significant expansion characteristics. The evolution process of gas migration channel and compaction area formed in the mining process of working face is shown in Fig. 2, and its dynamic evolution is closely related to the collapse process of overburden rock.
Fig. 2.
The law of crack propagation of the overlying rock in the working surface of the inclined extra-thick small coal column.
As shown in Fig. 2a, when the working face is mined for 32 m, the immediate roof of the overlying strata breaks and collapses to the floor of the coal seam. The caving heigh treaches 11.6 m, and a cavity with a height of 6.6 m is formed in the goaf. At the same time, four layers of separation cracks are formed in the overlying strata, and the height of these paration cracks reaches a maximum of 18 m from the floor of the coal seam.
As shown in Fig. 2a, when the working face is mined for 40 m, the first periodic weighting event occurs in the overburden rock, and the overall shape of the goaf changes greatly. The original four-layer separation cracks increase instantaneously and collapse occurs. The caving height increases instantaneously to 18.6 m, and a cavity with a height of 6.4 m is formed in the goaf. At this time, a hinged structure is formed for the first time in the goaf, and a new single-layer separation crack is formed at a distance of 21 m from the coal seam floor. As shown in Fig. 2a, when the working face advances to 45 m, the redistribution of stress in the overlying strata causes the second periodic weighting phenomenon, and the caving height reaches 21 m. The cavity with a height of 3 m is formed in the goaf, and 5 layers of separation cracks are formed in the overlying strata. The height of the separation cracks is up to 26 m from the coal seam floor.
As shown in Fig. 2a, when the working face is mined for 60 m, the third periodic weighting event occurs in the overburden rock, and the caving height increases instantaneously to 35.6 m. The cavity with a height of 1 m is formed in the goaf. The upper rock stratum develops under the action of stress and forms three layers of separation cracks. The maximum height of its development is 39 m from the roof of the coal seam. As shown in Fig. 2, when the working face is mined for 72 m, the fourth periodic weighting event occurs in the overlying strata, which causes a large area of overlying strata in the goaf to collapse, with the highest caving height reaching 46.8 m, and a cavity with a height of 0.5 m is formed in the goaf. At the same time, three layers of separation fissures are formed in the overlying strata, and the height of the separation fissures is up to 52 m from the coal seam floor.
As shown in Fig. 2a, when the working face is mined for 85 m, the fifth periodic weighting event occurs in the overburden rock. At this time, the caving speed of the overburden rock slows down, the caving height increases to 59.6 m, and a cavity of 0.5 m is formed in the goaf. Under the action of stress, the upper strata develop and form four layers of separation cracks, and the maximum height is 66.5 m from the roof of the coal seam. The final caving shape of the overburden rock is shown in Fig. 2b.
Through the analysis of Fig. 3, it is concluded that with the continuous advancement of the working face, the stress concentration factor in the stress concentration area further increases, and the maximum stress concentration factor in the stress concentration area increases from 1.5754 to 1.5978. When the fifth periodic weighting occurs, the stress concentration factor on the side of the working face continues to increase, and due to the collapse of the overlying strata in the goaf, the floor stress in the mining pressure relief area rises back to 0.2. At this time, the stress concentration coefficient of the bottom plate in the stress concentration area reaches a maximum of 1.6754.
Fig. 3.
The relationship between overlying rock movement and base plate stress evolution during recovery.
In view of the fact that the “O-X” overburden fracture structure formed by the coal seam trend in the actual mining has a leading role in the fracture trend, and the shearer is a continuous cutting operation, the tendency model adopts a one-step excavation method to simplify the simulation process. In the concrete implementation, after curing, the model is mined layer by layer from the bottom end to the top end in stages. After each stage of excavation, 200 mm × 60 mm × 60 mm pine strips are immediately placed to simulate the temporary support of hydraulic support. The quasi-static stress transfer path is constructed by the arrangement of the strips, and the cumulative support density is 1.2 frames/m. A total of 17 pieces of wood are arranged. In order to adapt to the overall size of the model, the size of the wooden strip needs to adapt to the overall size of 2000 mm × 200 mm × 2000 mm (tendency model). The length of 200 mm can correspond to the actual 20 m mining section distance on site, and the section of 60 mm × 60 mm can simulate the stress area of the support body after coal seam mining. It is arranged from the lower end to the upper end along the tendency, which is consistent with the actual law that the overlying rock collapse at the upper end is earlier than that at the lower end and the rock block is easy to slip downward after the mining of the gently inclined coal seam. In order to avoid the collapse of the whole model caused by the one-time removal of the support, the wood bars are arranged in stages and removed in a timely manner. The arrangement of the wood bars covers the whole inclined mining range. The mined-out area formed after the extraction can simulate the shape of the goaf in the working face, which provides a spatial basis for the subsequent 'O-X' breaking of the overlying strata, and ensures that the evolution of mining-induced fractures is consistent with the actual law of the site. In order to simulate the real mining stress release process, the wood strips in the low stress area are preferentially extracted, and finally the wood strips in the high stress area are removed synchronously to realize the instantaneous pressure relief of the whole section of the gently inclined working face. The experimental results show that the overburden damage presents a three-stage evolution law of “microcrack initiation-delamination propagation-macroscopic fracture”, and finally forms an asymmetric fracture network.
From Fig. 4, it can be seen that there is a significant difference between the distribution characteristics of mining-induced fractures in the overlying strata formed after the excavation of the inclined working face and the horizontal coal seam. This difference is mainly due to the downward slip of the collapsed overlying strata caused by the dip angle of the coal seam. The vertical layered structure composed of caving zone (25 m high) and fractured zone (81.7 m high) can be observed above the goaf, and the two sides of the working face show asymmetric development characteristics. The width of the upper fracture zone is 26 m and the caving angle is 71°, while the width of the lower fracture zone is reduced to 18 m and the caving angle is reduced to 60°. There is a unique “V” shape undisturbed area directly above the small coal pillar. From the left side of the working face to the right side, the shape of the fracture zone-compaction zone-fracture zone is presented in turn. The reason for this phenomenon is due to the dual mechanical effects: on the one hand, the dip angle of the coal seam and the mining stress jointly drive the upper and middle broken rock blocks to slide downward; on the other hand, the compressive stress of the overlying strata in the middle of the working face leads to the re-compaction of the mining cracks in the lower part.
Fig. 4.
Evolution of the crack field of inclined mining overlying rock.
It can be seen from Fig. 5 that when the mining of the working face is completed, the stress of the overlying strata is redistributed and reaches the equilibrium state, and the stress concentration factor of each stress concentration area in the goaf changes significantly compared with the original rock stress. The main performance is that the stress concentration coefficient on the upper side, the lower side and the support side of the small coal pillar increases significantly, and the stress concentration degree on the lower side of the inclined coal seam is more significant than that of the horizontal coal seam. The stress concentration factor in the compacted area is significantly lower than that of the original rock stress, and the stress concentration factor increases slowly from the upper side to the lower side. The maximum stress concentration factors of the upper side and the lower side are 1.87 and 1.51, respectively. The stress concentration factor increases from 0.40 to 0.80 from the upper side to the lower side in the compaction zone. The main reason for this effect is that the working face is affected by the dip angle and coal thickness after mining. When the caving rock mass of the overlying strata falls from the roof to the floor, it is affected by the component force of gravity along the tendency direction, and will slide along the working face from the upper side to the lower side, resulting in the lower side of the working face filling more fully than the upper side and the middle area, and the formed stress arch plays a supporting role in the overlying strata. However, at the same time, there is a significant stress concentration phenomenon on the support side of the upper coal pillar. Due to the mining of the 1801 goaf, the caving rock blocks are accumulated on the side of the coal pillar, and the shear force generated is superimposed with the gravity of the overlying strata, which makes the support side of the coal pillar form a large stress.
Fig. 5.
Stress characteristics of the bottom plate of the coal seam.
Numerical simulation study on fracture evolution characteristics of overlying strata in small coal pillar working face of inclined extra-thick coal seam
Software introduction
The method of 3DEC discrete element numerical simulation test can solve the problems of high cost, complicated operation and complex material model of traditional test, and can accurately quantify the value, which plays a very important role in scientific research. The 3DEC numerical simulation method includes 3DEC solid modeling, built-in FISH language programming to extend the effectiveness of 3DEC, 3DEC joint/contact surface/structural unit, static analysis, fluid–solid coupling, nonlinear dynamic simulation, 3DEC post-processing. The 3DEC software has the ability to analyze the physical properties of various continuous media, and also has the ability to study the changing process of discontinuous media under various conditions. At present, it is a common method to use 3DEC to study the movement and deformation characteristics of overlying strata under three-dimensional conditions in the simulation of coal mining strata (Fig. 6).
Fig. 6.
3DEC numerical simulation software and modeling.
In order to obtain the movement law of overlying strata on the fully-mechanized caving face of small coal pillars in inclined extra-thick coal seams, and to obtain the stress evolution of mining overlying strata and the development characteristics of fracture network, 3DEC numerical simulation analysis was carried out. The excavation distance of the model is set to 800 m, and the excavation step distance is set to 50 m, so as to analyze the displacement of the overlying strata in the process of coal seam propulsion.
When setting the initial model, it is necessary to consider that the model does not slip outside the region during calculation, so the boundary conditions of the model need to be set. The length of the numerical calculation model is set to be 1000 m, the height is 400 m, the width of the working face is 500 m, and the dip angle is 23°. The left and right sides of the model are respectively set with 100 m protective coal pillars to prevent offset in the model calculation process. The selection of lithological parameters of rock strata and joints is shown in Table 1.
Table 1.
Mechanical parameter properties of rock strata and joints.
| Lithologic characters | Rock formation | Joint | |||||||||
|---|---|---|---|---|---|---|---|---|---|---|---|
| Density KN (m3) | Bulk modulus (MPa) | Shear modulus (MPa) | Angle of internal friction (°) | Cohesion (MPa) | Tensile strength (MPa) | Bulk modulus (MPa) | Shear modulus (MPa) | Cohesion (MPa) | Angle of internal friction (°) | Tensile strength (MPa) | |
| Sandy mudstone | 2640 | 42,600 | 22,200 | 34 | 1.38 | 1.6 | 360 | 100 | 0.03 | 15 | 0.07 |
| Coal | 1460 | 10,480 | 5500 | 20 | 0.72 | 0.6 | 561.7 | 110 | 0.08 | 28 | 0.02 |
| mudstone | 1600 | 20,900 | 11,800 | 20 | 0.80 | 0.5 | 265.4 | 90 | 0.04 | 20 | 0.05 |
| Silt sandstone | 2000 | 36,900 | 21,800 | 35 | 1.30 | 1 | 647.1 | 200 | 0.25 | 30 | 0.11 |
| Medium sandstone | 2060 | 30,200 | 19,700 | 31 | 2.27 | 2 | 412.4 | 1390 | 2.20 | 27 | 0.60 |
| Coarse sandstone | 2640 | 38,200 | 22,200 | 34 | 1.38 | 5.68 | 647.1 | 200 | 0.33 | 30 | 0.11 |
According to the numerical simulation experiment scheme, the excavation process of the model working face is calculated, the slices are extracted along the strike of the model working face and the variation law of the subsidence displacement of the mining overburden at different advancing distances is analyzed, as shown in Fig. 7.
Fig. 7.
The variation law of subsidence displacement when advancing 100–800 m.
According to Fig. 7, in the early stage of the advancement of the 11,002 working face, the displacement of the overlying strata is small due to the limited spatial scale of the goaf. With the dynamic evolution of the original rock stress field, the separation cracks and fracture cracks are successively generated inside the immediate roof, and the separation cracks and fracture cracks are connected to eventually lead to the overall collapse of the immediate roof. As the working face continues to advance, a large cavity structure is formed at this time, which causes the upper rock mass to lose effective support, and induces the continuous collapse of the overlying rock, resulting in a significant increase in the vertical expansion height of the mining influence range. At the same time, multiple sets of gradually developed separation cracks are formed in the upper rock strata.
When the working face advances to the middle stage, the overburden system enters the stage of structural reconstruction. At this time, the spatial form of the beam-arch composite structure initially appears, but the hinge point has not been completely closed, and there are still large voids in the goaf. As the working face continues to advance, the falling rock mass on the upper side is affected by asymmetric mining stress, and the downward side tends to slip, resulting in significant anisotropic characteristics of fracture development in this area. The force of gravity along the tendency component promotes the directional migration of the caving rock mass, and the compaction area is formed on the lower side due to the accumulation effect of the caving rock mass.
When the inclination direction of the 11,002 working face continues to advance to the middle and late stages, the development of cavities in the goaf leads to the continuous enhancement of mining disturbance, and the overlying strata continue to deform and destabilize, forming a hinged structure along the lower side. The vertical activity space of the overlying strata increases, and some strata accumulate when they collapse, forming an irregular gap that is not completely filled at the upper part of the caving zone, resulting in a decrease in the subsidence displacement of the overlying strata.
When the inclination direction of the 11,002 working face continues to advance to the later stage, the central part of the upper uncollapsed rock stratum is more obviously subjected to the concentrated stress under the influence of mining disturbance. When the concentrated stress and the gravity of the rock stratum exceed the compressive strength of the rock stratum, the rock stratum deforms and bends until it breaks. Due to the influence of the inclination angle factor, it is affected by the component force of the gravity along the upper side to the lower side, and slips to the lower side. The overall performance is that the displacement of the overlying rock near the lower side is larger. The caving form of the whole overlying rock in the goaf is asymmetrically distributed, and a stable area is formed on the upper side of the small coal pillar, and the overlying rock does not collapse.
Distribution range and evolution characteristics of “two zones” of overlying strata in small coal pillar working face of inclined extra-thick coal seam
Transformation of “two zones” in goaf of small coal pillar working face in inclined extra-thick coal seam
Through the analysis of the results of physical simulation experiments, it is found that in the process of advancing the small coal pillar working face in the inclined thick coal seam, the development of mining-induced fractures in the goaf presents the characteristics of “two trapezoidal platforms”, of which two trapezoidal platforms correspond to the distribution areas of caving zone and fracture zone respectively. The three-dimensional spatial analysis shows that with the continuous advancement of the working face, the distribution range and dynamic evolution law of the caving zone and the fractured zone are clearly shown in Fig. 8. The model reveals the expansion process of mining-induced fractures from the three-dimensional dimension, and the spatial combination form of trapezoidal platform intuitively characterizes the variation characteristics of the two zones in different mining stages.
Fig. 8.
Distribution range and evolution law of falling zone and fissure zone.
The development mechanism of “two zones”: when the advancing distance of the working face reaches 25 m, the immediate roof breaks for the first time and collapses to fill the goaf, thus forming a caving zone with an initial height of 10.3 m. At this stage, the main roof only presents flexural deformation and is accompanied by the formation of separation cracks, and a complete fracture zone has not yet been formed. With the advance to 37 m, the highest caving height of overburden rock reaches 17.5 m, and the partition of caving zone and fracture zone appears for the first time. The height of caving zone is 12.5 m from the floor of coal seam, and the height of fracture zone is 12.5–17.5 m from the floor of coal seam.
The dynamic evolution of “two zones”: when the advancing distance reaches 45 m, the structural fracture of the middle sub-key stratum occurs, driving the penetrating fracture to extend to the upper sub-key stratum. After the key stratum is unstable, the rock mass at the bottom of the underlying fracture zone collapses structurally and enters the goaf, causing the local area of the original fracture zone to transform into the caving zone. In this stage, the vertical development height of the caving zone is increased to 20 m, and the fracture zone is readjusted to 20–35 m above the floor.
The spatial expansion law of “two zones”: the periodic collapse of overburden rock makes the range of two zones show a ladder-like expansion mode. When advances to 130 m, the overburden failure range and fracture development height reach a dynamic equilibrium state. The final vertical development height of the caving zone is stable at 25 m, and the fracture zone develops to the range of 25.2–81.7 m above the floor. The distribution of the two belts shows significant spatial heterogeneity. The horizontal extension of the propulsion direction becomes the dominant expansion mode, and the vertical development is gradually weakened.
Theoretical value calculation of mining overburden “two zones” height
Empirical calculation of caving zone height Empirical calculation formula of caving zone height is shown in Formula (1):
 |
1 |
In the formula: Hm is the height of the caving zone, m; M is the mining height of coal seam, m; Δ is the filling thickness caused by coal loss, m; kp is the bulking coefficient of caving rock; c is the total recovery rate, %; km is the bulking coefficient of caving top coal.
Based on the geological report of 11,002 working face, the mining height of the working face is 9.2 m, the broken expansion coefficient of caving top coal is 1.1, the broken expansion coefficient of caving rock is 1.3, and the total recovery rate is 85%. Substituting the above parameters into Formula (1), the maximum vertical height of the caving zone is calculated to be 25.60 m.
The empirical calculation formula of fracture zone height is shown in Table 2.
Table 2.
Experience formula for calculating the height of the crack zone.
| Lithology of overlying rock | Empirical formula for height calculation of caving zone |
|---|---|
| Hard lithology |  |
| Medium-hard lithology |  |
| Weak lithology |  |
| Extremely weak lithology |  |
According to the geological report of 11,002 working face, the lithology of roof overburden of B10 coal seam is hard overburden, and the maximum vertical height of fracture zone is 76.15 m based on the calculation formula of hard overburden.
Height determination of caving zone and fractured zone based on multi-method weight
In order to verify the accuracy of the development height of the caving zone and the fracture zone obtained by the evolution law of the overburden fracture, the theoretical values of the height of the two zones are obtained by the empirical formula for calculating the height of the caving zone and the fracture zone.
The principle of weighted average method
The weighted average method combines the advantages of each method and weakens its limitations by assigning the corresponding weight coefficients to the results of different methods, and finally obtains a more reliable prediction value. The mathematical expression is:
 |
2 |
In the formula: Hsynthesis is the comprehensive prediction height, m; wi is the weight percentage of the i th method. (∑wi = 1); Hi is the predicted value of the i th method, m。
“Two-band” height weight distribution calculation
The core principle of weight distribution is that the smaller the prediction error, the higher the reliability, and the greater the weight should be given. The specific calculation process is as follows:
In this study, the prediction data of caving zone height and fracture zone height listed in Table 3 are used as the basis of analysis.
Data pre-processing:
Table 3.
Height prediction value of caving zone and fracture zone.
| Prediction technique | Height of caving zone (m) | Fracture zone height (m) |
|---|---|---|
| Empirical formula | 25.05 | 64.95–76.15 |
| Method of physical simulation | 25.00 | 81.70 |
| Numerical simulation method | 23.50 | 94.21 |
The range value processing of the empirical formula of the fracture zone: The height of the fracture zone predicted by the empirical formula method is a range value (64.95–76.15 m). In order to facilitate the unified calculation error, the median value of the range is used as its representative value: the calculated empirical formula predicts that the height of the fracture zone is 70.55 m.
-
(2)
Reference value construction:
Due to the lack of measured truth values as absolute standards, this study constructs a “consensus reference value” as a benchmark for error calculation. The reference value is the arithmetic mean of the predicted values of the three methods: the calculated height of the reference caving zone is 24.52 m, and the height of the fractured zone is 82.15 m.
The reference value reflects the collective center position of the prediction results under the current data set, which is used to measure the degree of deviation of each method from the center.
-
(3)
Absolute error calculation:
Absolute error represents the absolute deviation between the predicted value and the reference value:
 |
3 |
-
(4)
Relative error calculation:
Due to the significant difference in the numerical scale of the height of the caving zone and the fracture zone, the relative error is standardized to eliminate the dimensional influence and make the errors of different zones comparable:
 |
4 |
-
(5)
The average relative error of the method is calculated:
In order to comprehensively evaluate the overall performance of each method in the prediction of caving zone and fracture zone, the average relative error across the two zones is calculated:
-
(6)Weight factor calculation:
5
The weight factor is inversely proportional to the square of the average relative error (inverse variance weighting idea), emphasizing the importance of the small error method:
 |
6 |
-
(7)
Normalized weight calculation:
The weight factor is normalized to obtain the final weight of each method, ensuring that the sum of ownership weights is 1:
 |
7 |
-
(8)
Calculation results:
According to the above methods, the prediction error and final weight of each method are calculated, and the results are summarized in Tables 4 and 5.
Table 4.
Prediction error calculation results of each method.
| Prognostic belt | Prediction technique | Predicted value (m) | Absolute error (m) | Relative error (%) |
|---|---|---|---|---|
| Fall band | Empirical formula | 25.05 | 0.53 | 0.0216 |
| (H reference value caving zone = 24. 52) | Method of physical simulation | 25.00 | 0.48 | 0.0196 |
| Numerical simulation method | 23.50 | 1.02 | 0.0416 | |
| Fracturation zone | Empirical formula | 70.55 | 11.60 | 0.1412 |
| Numerical simulation method | 94.21 | 12.06 | 0.1468 |
Table 5.
The average relative error and weight distribution results of the method.
| Prediction technique | Mean relative deviation | Weight factor | Normalization weight method | Weight percentage |
|---|---|---|---|---|
| Empirical formula | 0.0814 | 150.92 | 0.0228 | 2.28 |
| Method of physical simulation | 0.01255 | 6349.15 | 0.9603 | 96.03 |
| Numerical simulation method | 0.0942 | 112.68 | 0.0170 | 1.70 |
Height weighted calculation of caving zone and fractured zone
Through empirical formula, physical simulation and numerical simulation, the height of caving zone is 25.05 m, 25.00 m and 23.50 m respectively. The height of the fracture zone is 64.95–76.15 m, 81.70 m and 94.21 m, respectively. In order to integrate the advantages of multiple methods, it is necessary to assign weight percentages to various methods, as shown in Tables 6 and 7 below.
Table 6.
Weight distribution of caving zone.
| Method | Predicted value (m) | Weight percentage | Basis of distribution |
|---|---|---|---|
| Empirical formula | 25.05 | 2.28 | The statistical law is clear and the error is stable |
| Physical simulation | 25.00 | 96.03 | Intuitive reflection of the actual rupture process |
| Numerical simulation | 23.50 | 1.70 | The parameter sensitivity is high and the discreteness is large |
Table 7.
Weight distribution of fracture zone.
| Method | Predicted value (m) | Weight percentage | Basis of distribution |
|---|---|---|---|
| Empirical formula | 70.55 | 2.28 | Conservative design, avoid underestimating the risk |
| Physical simulation | 81.70 | 96.03 | The laboratory is highly controllable and the results are credible |
| Numerical simulation | 94.21 | 1.70 | Reflect extreme conditions, the need for risk early warning |
The high weight of physical similarity simulation is because the physical mechanism of real mining is directly reproduced. In the simulation process, the dynamic processes such as overburden collapse, separation fracture development and stress transfer can be observed in real time, and the physical phenomena such as asymmetric fracture expansion and rock block slip accumulation which are difficult to reproduce by numerical simulation can be captured. The multi-parameter synchronous measurement has high data credibility, which avoids the statistical deviation of empirical formula and the parameter assumption error of numerical simulation. The verification is highly consistent with the field engineering verification. Using the weighted average method based on weight, combined with the results of empirical formula, physical simulation and numerical simulation, the height of caving zone is 24.98 m and the height of fracture zone is 81.67 m calculated by Eq. (3).
Engineering practice of extraction technology in high permeability area of small coal pillar working face in inclined thick coal seam
Mining pressure relief gas extraction scheme of small coal pillar working face in inclined extra-thick coal seam
Drilling field layout parameters
Drilling field size: 3.6 m × 4 m × 2.6 m (width × depth × height); drilling spacing: 102 m.
-
(2)
Drilling layout parameters
The opening position: the upper and lower rows are arranged, the upper row of holes is 2.0 m from the bottom plate of the drilling field, and the lower row of holes is 1.5 m from the bottom plate of the drilling field. Ten high-level boreholes are arranged in the drilling field, and five high-level boreholes are arranged in parallel in the upper and lower rows. The lower rows of boreholes are numbered 1–1#, 1–2#, 1–3#, 1–4#, and 1–5#, respectively. The upper rows of boreholes are numbered 2–1#, 2–2#, 2–3#, 2–4#, and 2–5#, respectively. Among them, 1–1# and 2–1# boreholes are 0.4 m and 0.8 m away from the edge of the drilling field adit, respectively, and the remaining borehole spacing is 0.8 m. Design parameters of high-level borehole in drilling field of 11,002 working face as shown in Table 8 below.
Table 8.
Design parameters of high-level borehole in drilling field of 11,002 working face.
| Hole number | Hole diameter Φ (mm) | Angle with the wind lane (°) | Drilling elevation angle (°) | Final hole point (Distance from bottom plate) (m) | Deep hole (m) |
|---|---|---|---|---|---|
| 1–1# (Lower row) | 133 | Left avertence1.0 | 3.5 | 8 | 115.4 |
| 1–2# (Lower row) | 133 | Right deviation1.0 | 4.5 | 10 | 115.7 |
| 1–3# (Lower row) | 133 | Right deviation2.6 | 5.9 | 13 | 116.6 |
| 1–4# (Lower row) | 133 | Right deviation4.3 | 5.5 | 11 | 116 |
| 1–5# (Lower row) | 133 | Right deviation6.1 | 3.6 | 9 | 115.9 |
| 2–1# (Upper row) | 133 | 0 | 3.7 | 9 | 115.2 |
| 2–2# (Upper row) | 133 | Right deviation2 | 5.2 | 12 | 116.5 |
| 2–3# (Upper row) | 133 | Right deviation3.7 | 6.2 | 14 | 116.9 |
| 2–4# (Upper row) | 133 | Right deviation5.5 | 4.7 | 13 | 116.1 |
| 2–5# (Upper row) | 133 | Right deviation7.2 | 4.2 | 10 | 115.7 |
According to the actual situation of drilling in Tengda Coal Mine, in order to avoid the fracture of the connection of high-level boreholes, it is necessary to determine the optimal distance of drilling stubble. According to the actual situation of drilling in Tengda Coal Mine, it is determined that the stubble distance between 1# and 2# high-level drilling fields is 54 m, the stubble distance between 2# and 3# high-level drilling fields is 48 m, and the stubble distance between 3# and 4# high-level drilling fields is 50 m. as shown in Fig. 9.
Fig. 9.
High-level borehole layout diagram.
Analysis on gas extraction effect of small coal pillar working face in inclined extra-thick coal seam
Analysis of daily gas extraction amount in high drilling field
It can be seen from Fig. 10 that based on the continuous monitoring data of the 11,002 working face, the gas drainage volume shows a low and fluctuating state in the initial stage of mining. However, with the continuous advancement of the mining process, due to the disturbance effect caused by the activity of the working face, a gas migration channel is formed above the goaf. This change promotes a significant increase in gas drainage and gradually transitions to a relatively stable stage.
Fig. 10.
The relationship between absolute gas emission and gas extraction and air displacement.
During the mining period of the 11,002 small coal pillar working face, the monitoring data show that the absolute gas emission fluctuates between 22.12 and 39.21 m3 min−1, and the average value is 35.22 m3 min−1. According to statistical analysis, the range of air exhaust gas volume is 3.43–15.25 m3 min−1 (mean 7.24 m3 min−1), while the total gas drainage covers the range of 9.15–36.53 m3 min−1, and the average drainage efficiency is stable at 26.2 m3 min−1 level.
Figure 11 engineering data show that in the monitoring period from July to September 2023, the working face has completed 241.9 m footage and produced 385,000 tons of coal. Figure 12 Engineering data show that the gas concentration in key areas (working face, upper corner and return air roadway) is strictly controlled below the 1% threshold, which effectively guarantees the realization of the safe and efficient production goal of the fully mechanized caving face.
Fig. 11.
Daily output and cumulative footage change chart.
Fig. 12.
Change of gas concentration in upper corner and return airway.
Conclusion
The physical similarity simulation experiment of tendency and strike is designed, and the numerical simulation model of 11,002 working face is established. During the experiment, a total of 5 cycles of pressure were passed. In two stages (11–42 m from the coal pillar in the strike and 48–68 m from the coal pillar in the tendency); the separation cracks are fully developed in the range of 113–138 m from the coal pillar and 164–190 m from the coal pillar. With the increase of height, the surface fracture rate of overburden deformation and collapse shows a decreasing trend, and the change trend shows that the complexity of overburden fracture network decreases gradually from the working face.
Through the physical simulation and theoretical analysis of the development of mining-induced fractures in inclined extra-thick coal seams, the study reveals that the spatial distribution characteristics of overburden fractures are “two trapezoidal platforms”, corresponding to the caving zone and the fracture zone respectively. With the working face advancing to 130 m, the height of the two zones tends to be stable, the caving zone is finally 25 m, and the fracture zone is 25.2–81.7 m. Combined with the empirical formula and the weighted average method, it is verified that the height of the caving zone is 24.98 m and the height of the fracture zone is 81.67 m. The stepped expansion law and spatial differentiation characteristics are clarified, which provides a basis for safe mining of coal seams.
In this paper, the 11,002 small coal pillar working face of Tengda Coal Mine in Heze is taken as the prototype. Based on the combination of two-dimensional physical similarity simulation and numerical simulation, the evolution mechanism of mining-induced pressure-relieved gas high permeability zone in small coal pillar working face of inclined extra-thick coal seam is studied. By clarifying the evolution law of mining-induced overburden fracture in small coal pillar working face of inclined extra-thick coal seam, the temporal and spatial evolution characteristics of mining-induced fracture field and stress field morphology in small coal pillar working face of inclined extra-thick coal seam are further improved. The evolution law of pressure-relieved gas high permeability zone in small coal pillar working face of inclined extra-thick coal seam is mastered, and the spatial position of high permeability zone in small coal pillar working face of inclined extra-thick coal seam is accurately distinguished. It provides certain theoretical support for guiding the gas control work in the mine site, so that the gas concentration in the upper corner and return airway in the underground production process has been within the safety threshold range, which provides safety guarantee for the main mining face.
Author contributions
Conceptualization, Weidong Lu; Data curation, Pengxiang Zhao and Quan Jin; Formal analysis, Huan Jin; Funding acquisition, Shugang Li; Investigation, Jinyang Bai; Methodology, Shugang Li; Project administration, Shugang Li; Software, Weidong Lu; Supervision, Huan Jin; Visualization, Quan Jin; Writing—original draft, Quan Jin; Writing—review and editing, Pengxiang Zhao.
Funding
This research is supported by the Xinjiang Uygur Autonomous Region “Tianshan Talents” Plan (Leading Talents in Scientific and Technological Innovation) (2024TSYCLJ0026); General Project of National Natural Science Foundation of China (52174205); Key R & D task special project of Xinjiang Uygur Autonomous Region (2022B01034-3).
Data availability
The datasets used and/or analysed during the current study available from the corresponding author on reasonable request.
Declarations
Competing interests
The authors declare no competing interests.
Footnotes
Publisher’s note
Springer Nature remains neutral with regard to jurisdictional claims in published maps and institutional affiliations.
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Associated Data
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Data Availability Statement
The datasets used and/or analysed during the current study available from the corresponding author on reasonable request.