Hard roof-gas cooperative control method for fully mechanized caving mining of extra-thick coal seam

Abstract

The invention relates to the technical field of coal seam fully-mechanized caving mining, and discloses an extra-thick coal seam fully-mechanized caving mining hard roof-gas cooperative control method which comprises the following specific steps: collecting coal mine site stratum composition data; a numerical simulation physical model of the coal mine working face is established based on FLAC3D software, and the horizon of the hydraulic fracturing hard roof is obtained through simulation; mining-induced stress field inversion is carried out, and the fracturing implementation time and the drilling position are determined; according to the super-thick coal seam fully-mechanized caving mining hard roof-gas cooperative control method, the layer position of the hydraulic fracturing hard roof, the fracturing implementation time and the drilling position are obtained through simulation analysis, the position of a drilling end hole point is determined according to the goaf gas migration rule, and the gas migration rule of a mining field is determined by combining the expansion and evolution rule of hydraulic cracks and the mining field gas migration rule. The strong mine pressure weakening mechanism and the gas seepage path and gas pressure distribution characteristics in the gas extraction process are mastered, and effective control over the hard roof and efficient gas treatment are achieved.

Description

Technical Field

[0001] The invention relates to the technical field of fully-mechanized caving mining of coal seams, and in particular to a hard roof-gas coordinated control method for fully-mechanized caving mining of extra-thick coal seams. Background Art

[0002] Energy forms the basis of all production activities and daily life, and the sustained growth of the national economy depends on the growth of energy supply. China is both a major energy mining country and a major energy consumer, and its energy consumption accounts for more than 10% of the global total consumption. Coal, as a vital basic energy source in China, still accounts for 58% of the energy consumption structure during the planning period, and this proportion will remain stable for a long time in the future. Therefore, coal will continue to occupy a core position in China's energy strategy.

[0003] The geological conditions of China's coal seams are extremely complex. About one-third of the coal seams are covered with hard roofs that are difficult to collapse, and these coal seams are widely distributed in more than half of the mining areas in the country. In particular, in the 14 large-scale coal bases that are key to national construction, most mining areas are faced with the problem of hard roofs that are difficult to collapse. When there is such a hard roof above the coal seam, the crack spacing of the roof increases, causing the mining pressure in the mining area to be particularly severe. Especially when mining extra-thick coal seams, due to the large mining area space and the wide range of overburden movement, the movement law of the roof becomes more complicated and the mining pressure becomes more intense. Coal mines often encounter problems such as support crushing and tunnel damage. Field tests have shown that these problems can no longer be effectively solved by simply increasing the support strength.

[0004] In addition, during the high-intensity fully-mechanized caving process, the large amount of coal falling instantly causes a large amount of gas to gush out instantly. At the same time, the disturbance of the goaf by top coal caving will also cause gas to overflow from the goaf. These problems have caused gas to exceed the limit for a long time in the working face, return air lane, upper corner and rear tail during the mining process, which seriously threatens the safe and efficient production of the mine. However, the current treatment methods for hard roof and gas disasters often only target a single disaster factor, and do not combine the two for comprehensive treatment, and fail to achieve the coordinated control of hard roof and gas in fully-mechanized caving of extra-thick coal seams. This approach of neglecting one thing while focusing on the other leads to inefficient disaster management, long cycles and high costs. Summary of the invention

[0005] In order to solve the above problems that the current treatment methods for hard roof and gas disasters often only target a single disaster factor, do not combine the two for comprehensive treatment, fail to achieve coordinated control of hard roof and gas in fully-mechanized caving mining of extra-thick coal seams, and have low disaster treatment efficiency, long cycle and high cost, the present invention is implemented through the following technical solutions: A method for coordinated control of hard roof and gas in fully-mechanized caving mining of extra-thick coal seams, including the following specific methods:

[0006] Collecting data on the composition of the strata at the coal mine site;

[0007] A numerical simulation physical model of the coal mine working face was established based on FLAC3D software to simulate the hard roof layer for hydraulic fracturing;

[0008] Inversion of mining stress field to determine the timing and drilling location for fracturing;

[0009] Conduct hydraulic fracturing on the hard roof of large-space mining areas;

[0010] Determine the location of the drilling end point according to the gas migration law in the goaf;

[0011] Drill holes in the mining area to obtain surface fracturing wells, which also serve as extraction wells.

[0012] Furthermore, the collection of the on-site stratum composition data of the coal mine is specifically as follows:

[0013] Collect the composition and physical and mechanical parameters of the strata at the coal mine working face.

[0014] Furthermore, the numerical simulation physical model of the coal mine working face is a two-dimensional model, using weak interlayers to simulate the interface between rock layers, the top boundary of the model uses a free boundary, the bottom boundary uses a fixed boundary, and the horizontal displacement of the left and right sides of the model is restricted;

[0015] Through the numerical simulation physical model of the coal mine working face, each layer of hard roof is weakened respectively to simulate the change of the peak value of the working face advance support stress after the hard roof is weakened by hydraulic fracturing, and the relationship between the energy released by the destruction of each hard roof during longwall mining and the peak value of the working face advance support pressure is analyzed to determine the impact of the increase in the hard roof layer on the working face, so as to obtain the optimal hard roof layer for hydraulic fracturing.

[0016] Furthermore, through the mining stress field inversion, the stress distribution in the hard roof layer during the fracturing process is analyzed and simulated, and the expansion direction of the hydraulic fracture is predicted in combination with the expansion law of the hydraulic fracture. According to the expansion form of the hydraulic fracture, the timing and drilling position of the hydraulic fracture are determined, specifically:

[0017] Through numerical simulation methods, based on the mine pressure data during the actual mining process, the distribution of the mining stress field is inverted, and then the timing and drilling location of fracturing are determined based on the obtained data analysis.

[0018] Furthermore, based on the analysis of the inversion data of the mining stress field and guided by the expansion morphology of hydraulic fractures, the technical parameters of ground fracturing in the hard roof of large-space mining areas were optimized to obtain the optimal fracturing flow, fracturing pressure and fracturing duration.

[0019] Furthermore, the hard roof surface hydraulic fracturing is constructed by acidification and hydraulic sand addition composite fracturing, specifically:

[0020] Adding soil acid, which is a mixture of hydrochloric acid and hydrofluoric acid, to the pre-fluid, the hydrochloric acid reacts with the cementing material in the hard roof to weaken the cementation in the hard roof, and the hydrofluoric acid reacts with the silicate in the hard roof to weaken the skeleton particles in the hard roof;

[0021] After the reaction of the pre-fluid is completed, the hard roof is injected with fracturing fluid for hydraulic fracturing. The hard roof is fractured by hydraulic fracturing to reduce or eliminate the stress concentration phenomenon of the hard roof. Fine sand is added to the fracturing fluid as a proppant.

[0022] Furthermore, it also includes: using microseismic technology to monitor the dynamic expansion process of hydraulic fractures in real time, specifically:

[0023] Microseismic monitoring equipment is used to conduct on-site monitoring of multi-stage fracturing, and real-time monitoring of microseismic signals during hydraulic fracturing is performed to obtain the initiation location, expansion direction and geometric shape of hydraulic fractures.

[0024] Furthermore, the location of the drilling end point is determined according to the gas migration law of the goaf, specifically:

[0025] The mathematical model of seepage diffusion of gas migration in goaf uses the diffusion equation, gas outflow sink and basic differential equation of seepage as reference functions. According to Bernoulli equation and Darcy theorem, the orifice flow rate of the final hole of the drilling is:

[0026]

[0027] q is the orifice flow rate, in m / s; μA is the resistance coefficient of the goaf nozzle; g is the gravity acceleration, in m / s 2 ;z A 、z B are the height of the goaf and the height of the hole respectively; p A is the pressure of the goaf, p B is the pressure at the orifice, λ is g(ρ A -ρ B ), where ρ Ais the gas concentration in the goaf, ρ B is the gas concentration at the orifice; k is the permeability coefficient; h is the height, in meters, and s is the length of the seepage path;

[0028] The diffusion movement of coal seam gas in the goaf follows the diffusion equation in the three-dimensional porous medium, that is, the gas outflow differential equation group, which is:

[0029]

[0030] c is the volume percentage concentration of coal seam gas in the goaf; t is time, the unit is s; is the Hamiltonian operator; is the diffusion coefficient; is the wind flow seepage velocity in the goaf; c' is the gas concentration released by the falling gangue; W is the gas distribution source intensity;

[0031] Substituting the outflow source into the basic differential equations of seepage, we get:

[0032]

[0033] Among them, a is ν is the kinematic viscosity coefficient of the mixed gas, the unit is m 2 / s, g is the acceleration due to gravity, the unit is m / s 2 ; b is β is the geometric shape factor, n is the porosity, d is the medium equivalent diameter, the unit is m, e is the permeability scalar coefficient; p is the static pressure; z is the height of the goaf, the unit is m; h 静 is the hydrostatic head; B is the dimension correction coefficient; φ is the porosity in the coal rock mass; C is the isothermal adsorption constant; Where e is a constant, b0 is the decay rate, t0 is the initial time, t1 is the decay time; H is the caving height of the goaf;

[0034] The porosity of the coal rock mass is assumed to be constant over time and pressure, and the differential equations of the orifice flow rate of the final hole of the drilling, the gas outflow sink, and the basic differential equations of the seepage based on the outflow source are combined to obtain the boundary conditions. Based on the numerical solution, the law of the change of gas concentration in the goaf over time is obtained.

[0035] According to the judgment of the flow state by Reynolds number, under the original permeability, gas forms linear seepage in the coal seam, and the seepage displacement distance is L1:

[0036]

[0037] ρ is the fluid density; k is the permeability; Δp is the pressure difference; μ is the gas dynamic viscosity; Re is the Reynolds number;

[0038] When the gas flow rate is lower than the flow rate during linear seepage, there is a starting pressure gradient λ0 for the gas. When the pressure gradient is greater than the starting pressure gradient, the gas migration in the coal seam is nonlinear seepage, and the seepage displacement distance is L2:

[0039]

[0040] The gas extraction radius is the sum of the seepage displacement distance L1 and the seepage displacement distance L2.

[0041] Furthermore, the boundary conditions include:

[0042] The first boundary condition gives the pressure distribution of gas at the extraction wellhead;

[0043] The second boundary condition gives the velocity distribution of gas at the extraction wellhead;

[0044] The third boundary condition gives the relationship between the internal and external pressure difference on the boundary and the gas flow rate.

[0045] Compared with the prior art, the present invention has the following beneficial effects:

[0046] This method of coordinated control of hard roof and gas in fully-mechanized caving mining of extra-thick coal seams obtains the stratum position of the hard roof for hydraulic fracturing, the timing of fracturing and the drilling position through simulation analysis, and determines the position of the drilling end point according to the gas migration law in the goaf. Combined with the expansion and evolution law of hydraulic fractures and the migration law of gas in the mining area, it masters the weakening mechanism of strong mine pressure and the gas seepage path and distribution characteristics of gas pressure during gas extraction, thereby achieving effective control of the hard roof and efficient gas management. BRIEF DESCRIPTION OF THE DRAWINGS

[0047] Figure 1 It is a flow chart of a method for coordinated control of hard roof and gas in fully mechanized top caving mining of extra-thick coal seams in an embodiment of the present invention;

[0048] Figure 2 It is a schematic diagram of the peak value of the supporting pressure and gravity of each layer of hard roof of the working face acting on the coal seam in an embodiment of the present invention;

[0049] Figure 3 Schematic diagram of the curve of gas concentration in goaf changing with time in an embodiment of the present invention;

[0050] Figure 4 Schematic diagram of a gas extraction model for a vertical fracturing well in an embodiment of the present invention;

[0051] Figure 5 The cloud diagram of gas pressure distribution at different extraction times in the embodiment of the present invention;

[0052] Figure 6 This is a numerical calculation model diagram of mining stress evolution in an embodiment of the present invention;

[0053] Figure 7 This is a layout diagram of ground gas extraction drilling holes in an embodiment of the present invention;

[0054] Figure 8 It is a numerical model diagram of FLAC3D software in an embodiment of the present invention. DETAILED DESCRIPTION

[0055] The following will be combined with the drawings in the embodiments of the present invention to clearly and completely describe the technical solutions in the embodiments of the present invention. Obviously, the described embodiments are only part of the embodiments of the present invention, not all of the embodiments. Based on the embodiments of the present invention, all other embodiments obtained by ordinary technicians in this field without creative work are within the scope of protection of the present invention.

[0056] The embodiment of the hard roof-gas coordinated control method for fully-mechanized caving mining of extra-thick coal seams is as follows:

[0057] See also Figure 1-Figure 8 , a method for coordinated control of hard roof and gas in fully mechanized top-caving mining of extra-thick coal seams, including the following specific methods:

[0058] Collecting data on the composition of the strata at the coal mine site, specifically collecting the composition and physical and mechanical parameters of the strata at the coal mine working face;

[0059] A numerical simulation physical model of the coal mine working face was established based on FLAC3D software to simulate the hard roof layer for hydraulic fracturing;

[0060] Inversion of mining stress field to determine the timing and drilling location of hydraulic fracturing. Specifically, the distribution of mining stress field is inverted based on the ore pressure data in the actual mining process through numerical simulation method, and then the timing and drilling location of hydraulic fracturing are determined based on the obtained data analysis;

[0061] Optimize the technical parameters of ground fracturing in large-space mining areas with hard roof to obtain the optimal fracturing flow, fracturing pressure and fracturing duration;

[0062] Conduct hydraulic fracturing on the hard roof of large-space mining areas;

[0063] Use microseismic technology to monitor the dynamic expansion process of hydraulic fractures in real time;

[0064] Determine the location of the drilling end point according to the gas migration law in the goaf;

[0065] Drill holes in the mining area to obtain surface fracturing wells, which also serve as extraction wells.

[0066] Taking the 8101 working face of Tashan Coal Mine as the engineering background, the following specific implementation is carried out:

[0067] The hard roof of the coal mine is mainly composed of fine sandstone, medium sandstone, coarse sandstone, sandy mudstone, etc., with sandy rock accounting for 90%-95%. At the same time, the quartz content of the hard roof sandstone was measured to be 67.3%, clay minerals 18.6%, and the remaining components were mainly calcite, plagioclase and ankerite. There are multiple layers of hard roof, and the mine pressure is strong. The specific working face strata and their physical parameters are shown in Table 1:

[0068] Table 1

[0069]

[0070]

[0071]

[0072]

[0073] Based on FLAC3D software, a physical model of the 8101 working face of Tashan Coal Mine was established for numerical simulation. A numerical model containing five layers of hard roof was established, such as Figure 8 The model size is 1200mx580m (length x height), and the number of units is 36611. An extremely thin weak interlayer is used to simulate the interface between rock layers. At the same time, in order to avoid the reduction of calculation accuracy and rate by building a three-dimensional model of mine scale, a two-dimensional model is used. The top boundary of the model adopts a free boundary, the bottom boundary adopts a fixed boundary, and the horizontal displacement on the left and right sides is restricted.

[0074] The simulation of the rock plastic zone at the completion of mining and the rock migration at the completion of mining are quite consistent with the actual observations, indicating that the data is reasonable. The relationship between the energy released by the destruction of each hard roof during longwall mining and the peak value of the advance support pressure of the working face is analyzed. It shows that with the increase of the hard roof layer, the above corresponding relationship gradually weakens, that is, the impact of the increase in the hard roof layer on the working face is reduced. Specifically, each layer of the hard roof is weakened separately under the model to simulate the change in the peak value of the advance support stress of the working face after the hard roof is weakened by hydraulic fracturing. In summary, when the thickness of the far-field hard roof is not much different, the relatively low-level hard roof has a greater impact on the mine pressure of the working face. Therefore, when fracturing the far-field hard roof to control the strong mine pressure of the working face, the hard roof with the lowest layer should be fractured first. Such as Figure 2 As shown in the figure, with the 8101 working face of Tashan Coal Mine as the background, the best target layer for hydraulic fracturing is the lowest layer 1# hard roof, followed by the 2# hard roof above it. If these two layers of hard roof are fractured, the effect of alleviating the strong mine pressure on the working face will be better.

[0075] The results of the EH-4 electromagnetic imaging system show that the heights of the collapse zone and the fracture zone in the mining area are 100m and 170m respectively, the collapse angle of the overburden is about 65°, and the rotation angle of the masonry rock block is 10°. Combined with the parameters in Table 1, the peak value of the bearing pressure and gravity of each layer of hard roof acting on the coal seam are obtained, as shown in Figure 2 As shown, the analysis shows that the 60# roof (5# hard roof) of the 8101 working face of Tashan Coal Mine is the most pressurized layer, and the pressure on the coal seam of the working face is the greatest. Therefore, hydraulic fracturing of the hard roof of this layer can be considered. When only one layer of hard roof is fractured, the 34# roof (1# hard roof) is preferred; when two layers of roof are fractured, the 34# and 40# roofs (2# hard roofs) are selected. This is based on the two layers of far-field hard roofs that need to be fractured and are not in the curved sinking zone obtained from the analysis of the rock pressure of the mining area. In addition, there is no third layer of roof that should or is worth fracturing.

[0076] Through mining stress field inversion, the stress distribution in the hard roof layer during the fracturing process is analyzed and simulated. The expansion direction of the hydraulic fractures is predicted based on the expansion law of the hydraulic fractures. According to the expansion form of the hydraulic fractures, the timing and drilling location of the hydraulic fractures are determined. Specifically:

[0077] According to the engineering geological profile of the 8101 working face, the thinner rock layers were merged and simplified to establish a numerical calculation model diagram, such as Figure 6 As shown. The length (x) of the calculation model is 300m, the width (y) is 250m, and the height (z) is 166m. The distance between the coal seam and the bottom of the model is 7m, and the mining area is located in the middle of the coal seam. The size is 180m×120m×22m (x, y, z). The model has a total of 25 layers, and ultra-thin contact layers are established between the layers with a height of 0.1m. The model parameters are shown in Table 2:

[0078] Table 2

[0079]

[0080]

[0081] The model takes into account the gravity effect and sets the gravity acceleration to -9.8m / s 2 The upper part of the model adopts fixed stress constraint, that is, the rock layer between the upper part of the model and the surface is converted into gravity and applied to the upper boundary of the model, and the rest of the model boundaries adopt fixed displacement constraint.

[0082] The initial stress state of the coal mining face is: the maximum horizontal principal stress (x direction) is 12MPa, the minimum horizontal principal stress (y direction) is 6.4MPa, and the vertical principal stress (z direction) is 11.4MPa. Based on this stress state, the on-site geostress field is inverted. After the stress is balanced, the longwall mining process is simulated. The distance between the coal seam and the bottom of the model is 7m, the cut is located at x=60m, the mining range along the dip (y) is 65m≤y≤185m, and each step along the strike (z) is 5m. A total of 36 steps are mined until x=240m. After each mining, sufficient calculation steps (40000) are set to ensure that the stress of the surrounding rock around the goaf is fully released and the overburden is broken.

[0083] Two different fracturing position experiments were conducted: when in the middle of the goaf along the strike, water injection points were set at different working face advancement distances (20m, 60m, 100m, 140m, 200m) to perform hydraulic fracturing on the hard roof; when in front of the working face along the strike, the water injection positions were set at 20m, 40m, and 60m from the working face on the strike section; secondly, three water injection points were selected on the dip section, located at y=50m, 130m, and 200m respectively. Water was injected from the middle of the water injection point at an injection rate of 1m 3 / min, injection time is 30min.

[0084] It can be found that when the hydraulic fracture is in the middle of the goaf along the strike, affected by the mining stress, the expansion process of the hydraulic fracture is accompanied by a certain degree of deflection, forming an "arc-shaped" crack, and the turning angle increases with the continuous advancement of the working face. Among them, the hydraulic fracture deflection at the 140m position is the largest, gradually deflecting from vertical cracks to horizontal cracks; the hydraulic fracture deflection angle in the stress-increasing areas on both sides is relatively small. When the working face is in front of the strike, the high-position hard roof rock layer in front of the working face is less affected by the mining stress, and no significant deflection of the hydraulic fracture is induced.

[0085] The study shows that the influence of mining stress disturbance on different strata above the goaf is significantly different, so the stress state of different strata is different; generally speaking, the influence of mining stress disturbance gradually decreases with the increase of distance from the roof. In order to explore the morphological characteristics of mining stress on the expansion of hydraulic fractures in the hard roof, the 5th, 6th and 7th strata in Table 2 are combined and set as the hard roof, and the original hard roof stratum (21st stratum) is set as a non-critical layer. For the convenience of comparison, the water injection position is also set according to the mining stress state when the working face advances 140m: the water injection position is set at 20m, 40m and 60m from the working face on the strike section; secondly, three water injection points are selected on the dip section, located at y=50m, 130m and 200m respectively. It can be found that when the water injection position is 20m away from the working face along the strike, the deflection angle of the hydraulic fracture is the largest, especially the hydraulic fracture located at y=130m, which forms an approximately horizontal "arc-shaped" crack on the side close to the working face and a vertical crack on the side away from the working face. When the water injection position is more than 40m away from the working face along the strike, the morphology of the hydraulic fractures formed at each water injection point is not much different. By comparison, it can be seen that when hydraulic fracturing is performed on hard roofs at different levels, the closer the fracturing layer is to the mining site, the longer the path length experienced by the hydraulic fracture from the strong disturbance zone of mining stress to the weak disturbance zone, the greater the cumulative deflection angle of the hydraulic fracture, and the more complex the fracture network formed. Based on the above experience and in order to reduce the harm of gas to coal mine production, combined with the needs of underground coal mining planning in Tashan Mine, gas is extracted in conjunction with ground horizontal fracturing drilling, and comprehensive consideration of various factors is taken into account. Initially, 7 ground boreholes were arranged at the 8204 working face of Tashan Mine. See the specific location. Figure 7 .

[0086] By analyzing the data from the above-mentioned mining stress field inversion and taking the expansion morphology of hydraulic fractures as a guide, the technical parameters of ground fracturing on the hard roof of large-space mining areas are optimized to obtain the optimal fracturing flow, fracturing pressure and fracturing duration.

[0087] The technical equipment for weakening the hard roof by hydraulic fracturing on the ground mainly includes main pressure pump truck, acid pump truck, acid tank truck, sand tank truck, sand mixer truck, manifold truck and instrument truck. In order to weaken the hard roof of the working face and reduce the initial pressure strength of the working face; at the same time, the permeability of the top coal is enhanced to make the top coal collapse in time, avoid the concentrated release of gas and the large-scale gas outflow from the goaf at one time, so as to achieve the purpose of initial mining gas over-limit and strong mine pressure. The construction of weakening the hard roof by hydraulic fracturing on the ground is as follows:

[0088] 1. Use acidification and hydraulic sand fracturing to carry out construction;

[0089] Second, since the target hard roof of the fracturing is sandstone, mud acid (a mixture of hydrochloric acid and hydrofluoric acid) is added to the pre-fluid, and its main components are HCL and HF. HCL reacts with the cement in the sandstone to weaken the sandstone cementation, and HF reacts with the silicate in the sandstone to weaken the skeleton particles in the sandstone, thereby achieving chemical weakening of the hard roof;

[0090] 3. After the pre-acid injection is completed and the pump is stopped for reaction, a large amount of fracturing fluid is injected into the hard roof for hydraulic fracturing. Through hydraulic fracturing, cracks are created, the hard roof is cracked, the stress concentration phenomenon of the hard roof is reduced or eliminated, and the occurrence of impact ground pressure and strong mine pressure is effectively prevented through physical destruction.

[0091] It should be noted that in order to enhance the fracturing effect and help the extension of cracks, a certain amount of fine sand is added as a proppant during the fracturing construction. On the one hand, the use of fine sand can reduce liquid loss and improve liquid efficiency; on the other hand, the grinding of the crack wall by fine sand can effectively reduce the generation of curved cracks and facilitate the formation of long cracks.

[0092] In addition, in order to improve the effect of composite fracturing, the well is fully shut down after fracturing and no blowdown is carried out to ensure sufficient time for acid-rock reaction and fully weaken the hard roof.

[0093] Microseismic technology is used to monitor the dynamic expansion process of hydraulic fractures in real time during ground hydraulic fracturing. Hydraulic fracturing of rock formations will cause rock formation fractures, and at the same time, corresponding microseismic events will be generated. By monitoring the microseismic signals during the fracturing process and further diagnosing and processing them, the starting position, expansion direction and geometric morphology of the hydraulic fractures can be obtained. According to the design of the fracturing scheme, the first to fourth level fracturing field tests were carried out on the fracturing of the Tashan 8204-2 well. The starting position of the first level fracturing is located near the end point of the fracturing, and the artificial fractures extend in the NE40° azimuth as a whole, and the cracks on the east and west wings extend evenly; the starting position of the second level fracturing is located near the end point of the fracturing, and the artificial fractures extend in the NE54° azimuth as a whole; the starting position of the third level fracturing is located near the end point of the fracturing, the cracks on the west wing develop earlier, the east and west wings are roughly symmetrical, and the expansion of the west wing is better than that of the east wing; the starting position of the fourth level fracturing is located near the end point of the fracturing, the cracks on the west wing develop earlier, the east and west wings are roughly symmetrical, and the expansion of the west wing is better than that of the east wing. According to the analysis of the fracturing crack detection results of the entire well section, it can be seen that obvious effective cracks have been formed at all levels of fracturing, the cracks have roughly the same direction, and the east and west wings of each level of fracturing are basically symmetrical.

[0094] Taking the 76# support in the middle of the working face as an example, the support resistance and mine pressure manifestation characteristics in the non-fracturing affected area (210-355m) and the fracturing affected area (355-490m) were analyzed. The measurement points were arranged at intervals of 20m on the top and side of the tunnel within the range of 210-355m and 355-490m from the cutting eye of the working face to analyze the displacement changes of the tunnel: hydraulic fractures can reduce the periodic pressure intensity in the hydraulic fracture affected area. The maximum working resistance of the 76# support before fracturing was 44.3MPa, and the maximum working resistance after fracturing was reduced to 38.5MPa; the average working resistance dropped from 33.1MPa to 27.5MPa, a reduction of 16.9%. Hydraulic fractures reduce the periodic pressure step. After hydraulic fracturing of the high-level hard roof, the maximum pressure step and the average periodic pressure step are reduced. The average periodic pressure step is reduced by 15%, about 3.5m. Before fracturing, the mining position is near 280m and 305m, and strong mine pressure appears on the working face. The pressure lasts for 12m and 23m respectively. The #45-#90 support safety valve is opened, and the coal wall is severely spalled. After fracturing, in the area affected by hydraulic fractures, the intensity of mine pressure on the working face is significantly reduced, and no strong mine pressure appears on the working face. The opening rate of the support safety valve decreases by 37%. In the area affected by hydraulic fracturing, the deformation of the roadway is significantly improved, and the deformation of the roadway height is reduced from an average of 0.99m to an average of 0.53m, and the deformation of the roadway width is reduced from an average of 0.45m to an average of 0.23m. There is no damage to the single pillar in the roadway, and the advanced support section of the working face mining roadway is well controlled.

[0095] In addition, when hydraulic fracturing was not performed, the working face had strong mine pressure, macro cracks appeared in the coal wall and coal wall spalling, while after hydraulic fracturing, the mine pressure on the working face was significantly weakened, and the coal wall was intact without spalling. In summary, after hydraulic fracturing of the hard roof of the 8204-2 working face of Tashan Coal Mine on the ground, the mine pressure on the working face was significantly reduced, indicating that after hydraulic fracturing of the hard roof, a significant control effect of strong mine pressure was achieved, solving the problem of insufficient high-level hard roof that cannot be controlled in traditional underground mines, which is of great significance for solving similar mine pressure disasters induced by the instability of high-level hard roof.

[0096] Due to the existence of gaps between the falling gangue, there is a gas flow field inside the goaf, and the gas concentration in the goaf is not the same everywhere. Therefore, the location of the drilling end point directly affects the extraction effect of ground drilling. In order to achieve the optimal effect of ground drilling gas extraction, it is very important to determine the reasonable location of the drilling end point.

[0097] According to the gas migration law in the goaf, the location of the drilling end point is determined. The mathematical model of seepage diffusion of gas migration in the goaf uses the diffusion equation, gas outflow sink and seepage basic differential equation as reference functions. The pressure of the goaf is set as p A The pressure at the orifice of the final hole of the drilling is pB , the velocity in the goaf is v A The velocity at the orifice of the final hole is v B , the inlet loss of the orifice is introduced, and according to the Bernoulli equation and Darcy's theorem, the orifice flow rate of the final hole of the drilling is:

[0098]

[0099] q is the orifice flow rate, in m / s; μA is the resistance coefficient of the goaf nozzle; g is the gravity acceleration, in m / s 2 ;z A 、z B are the height of the goaf and the height of the hole respectively; λ is g(ρ A -ρ B ), where ρ A is the gas concentration in the goaf, ρ B is the gas concentration at the orifice; k is the permeability coefficient; h is the height, in meters, and s is the length of the seepage path;

[0100] The diffusion movement of coal seam gas in the goaf follows the diffusion equation in the three-dimensional porous medium, that is, the gas outflow differential equation group, which is:

[0101]

[0102] c is the volume percentage concentration of coal seam gas in the goaf; t is time, the unit is s; is the Hamiltonian operator; is the diffusion coefficient; is the wind flow seepage velocity in the goaf; c' is the gas concentration released by the falling gangue; W is the gas distribution source intensity;

[0103] Substituting the outflow source into the basic differential equations of seepage, we get:

[0104]

[0105] Among them, a is ν is the kinematic viscosity coefficient of the mixed gas, the unit is m 2 / s, g is the acceleration due to gravity, the unit is m / s 2 ; b is β is the geometric shape factor, n is the porosity, d is the medium equivalent diameter, the unit is m, e is the permeability scalar coefficient; p is the static pressure; z is the height of the goaf, the unit is m; h 静 is the hydrostatic head; B is the dimension correction coefficient; φ is the porosity in the coal rock mass; C is the isothermal adsorption constant; Where e is a constant, b0 is the decay rate, t0 is the initial time, t1 is the decay time; H is the caving height of the goaf.

[0106] Assuming that the porosity of the coal rock mass does not change with time and pressure, the orifice flow rate of the final hole of the drilling, the gas outflow sink differential equation group and the basic seepage differential equation group based on the outflow source are combined to obtain the boundary conditions. The first boundary condition gives the pressure distribution of the gas at the extraction wellhead; the second boundary condition gives the velocity distribution of the gas at the extraction wellhead; the third boundary condition gives the relationship between the internal and external pressure difference on the boundary and the gas flow rate.

[0107] After setting certain boundary conditions, the final drilling hole position is obtained according to the numerical solution, that is, according to the simulation of FLAC3D software, combined with the above three boundary conditions, and the law of gas concentration change in the goaf over time is obtained. The specific curve of gas concentration change in the goaf over time is as follows: Figure 3 shown.

[0108] According to the judgment of Reynolds number on flow state, it is difficult for gas flow in coal seams to form high-speed nonlinear seepage. Under the original seepage rate, gas forms linear seepage in coal seams, and the seepage displacement distance is L1:

[0109]

[0110] ρ is the fluid density; k is the permeability; Δp is the pressure difference; μ is the gas dynamic viscosity; Re is the Reynolds number, Re = 10 -4 It is the dividing line between linear seepage and nonlinear seepage.

[0111] Under lower speed conditions, that is, when the gas flow rate is lower than the flow rate during linear seepage, there is a starting pressure gradient λ0 for the gas. When the pressure gradient is greater than the starting pressure gradient, the gas migration in the coal seam is nonlinear seepage, and the seepage displacement distance is L2:

[0112]

[0113] The gas extraction radius is the sum of the seepage displacement distance L1 and the seepage displacement distance L2. According to the extraction related parameters, it is calculated that: L1 = 31.473m, L2 = 0.789m, then the effective extraction radius of the borehole is 32.262m.

[0114] The drilling holes in the mining area are the coal-bed methane extraction holes constructed after the formation is affected by coal mining, targeting the residual coal in the mining area and the gas stored in the surrounding rock. The gas source, geological conditions and development process are significantly different from the conventional ground coal-bed methane pre-extraction wells. Therefore, the following principles should be followed in the arrangement of drilling holes: 1. The drilling holes are easy to extract; 2. The drilling holes are not easily damaged by mining; 3. The hole spacing is reasonable; 4. The ground environment is favorable. Take the 8204-2 working face as an example, where the vertical well is located in the middle of the working face, consisting of a vertical section, a variable inclination section, and a horizontal section. The designed length is 650m, the vertical section is 120m, the inclination section is 330m, and the horizontal section is 200m. Aperture: 0-30m inner diameter is 444.5mm, 30-120m inner diameter is 311.5mm, 120-330m inner diameter is 216mm. The fracturing was carried out in three stages, with a total of three sections, namely, perforation fracturing in the borehole length sections of 573-623m, 503-526m, and 444-475m. After fracturing, the vertical well also served as a gas extraction well.

[0115] like Figure 4 and Figure 5 As shown in the figure, while the ground fracturing method is used to control the thick and hard roof rock layer in the far field to effectively control the strong mine pressure, the vertical well of the ground fracturing forms a vertical fracturing well and also serves as a drainage well. It can be clearly seen from the figure that in the early stage of drainage, the gas pressure value in the coal seam is relatively large. With the increase of drainage time, the gas pressure in the coal seam gradually decreases, and the gas concentration also decreases accordingly. Figure 5 (a), (b), (c), (d), (e), (f), (g), and (h) are the gas pressure values ​​on the 0th day, the 1st day, the 5th day, the 10th day, the 15th day, the 20th day, the 25th day, and the 30th day, respectively, in Pa, which proves that it can effectively extract gas and reduce gas concentration.

[0116] While the embodiments of the present invention have been shown and described, it will be appreciated by those skilled in the art that many changes, modifications, substitutions and variations can be made to the embodiments without departing from the principles and spirit of the invention.

Claims

1. A method for coordinated control of hard roof and gas in fully mechanized caving mining of extra-thick coal seams, characterized in that: The following specific methods are included: Collecting data on the composition of the strata at the coal mine site; A numerical simulation physical model of the coal mine working face was established based on FLAC3D software to simulate the hard roof layer for hydraulic fracturing; Inversion of mining stress field to determine the timing and drilling location for fracturing; Conduct hydraulic fracturing on the hard roof of large-space mining areas; Determine the location of the drilling end point according to the gas migration law in the goaf; Drill holes in the mining area to obtain surface fracturing wells, which also serve as extraction wells.

2. The method for coordinated control of hard roof and gas in fully mechanized top caving mining of extra-thick coal seams according to claim 1 is characterized by: The collected coal mine on-site stratum composition data specifically includes: Collect the composition and physical and mechanical parameters of the strata at the coal mine working face.

3. The method for coordinated control of hard roof and gas in fully mechanized top caving mining of extra-thick coal seams according to claim 1 is characterized by: The numerical simulation physical model of the coal mine working face is a two-dimensional model, which uses weak interlayers to simulate the interface between rock layers, the top boundary of the model uses a free boundary, the bottom boundary uses a fixed boundary, and the horizontal displacement of the left and right sides of the model is restricted; Through the numerical simulation physical model of the coal mine working face, each layer of hard roof is weakened respectively to simulate the change of the peak value of the working face advance support stress after the hard roof is weakened by hydraulic fracturing, and the relationship between the energy released by the destruction of each hard roof during longwall mining and the peak value of the working face advance support pressure is analyzed to determine the impact of the increase in the hard roof layer on the working face, so as to obtain the optimal hard roof layer for hydraulic fracturing.

4. The method for coordinated control of hard roof and gas in fully mechanized top caving mining of extra-thick coal seams according to claim 1 is characterized in that: Through the mining stress field inversion, the stress distribution in the hard roof layer during the simulated mining process is analyzed, and the expansion direction of the hydraulic fracture is predicted in combination with the expansion law of the hydraulic fracture. According to the expansion form of the hydraulic fracture, the timing and drilling position of the hydraulic fracture are determined, specifically: Through numerical simulation methods, based on the mine pressure data during the actual mining process, the distribution of the mining stress field is inverted, and then the timing and drilling location of fracturing are determined based on the obtained data analysis.

5. The method for coordinated control of hard roof and gas in fully mechanized top caving mining of extra-thick coal seams according to claim 4 is characterized in that: Based on the analysis of the inversion data of the mining stress field and guided by the expansion morphology of hydraulic fractures, the technical parameters of ground fracturing on the hard roof of large-space mining areas are optimized to obtain the optimal fracturing flow, fracturing pressure and fracturing duration.

6. The method for coordinated control of hard roof and gas in fully mechanized top caving mining of extra-thick coal seams according to claim 5 is characterized by: The hard roof surface hydraulic fracturing is constructed by acidification and hydraulic sand addition composite fracturing, specifically: Adding soil acid, which is a mixture of hydrochloric acid and hydrofluoric acid, to the pre-fluid, the hydrochloric acid reacts with the cementing material in the hard roof to weaken the cementation in the hard roof, and the hydrofluoric acid reacts with the silicate in the hard roof to weaken the skeleton particles in the hard roof; After the reaction of the pre-fluid is completed, the hard roof is injected with fracturing fluid for hydraulic fracturing. The hard roof is fractured by hydraulic fracturing to reduce or eliminate the stress concentration phenomenon of the hard roof. Fine sand is added to the fracturing fluid as a proppant.

7. The method for coordinated control of hard roof and gas in fully mechanized top caving mining of extra-thick coal seams according to claim 6 is characterized in that: Also includes: Microseismic technology is used to monitor the dynamic expansion process of hydraulic fractures in real time, specifically: Microseismic monitoring equipment is used to conduct on-site monitoring of multi-stage fracturing, and real-time monitoring of microseismic signals during hydraulic fracturing is performed to obtain the initiation location, expansion direction and geometric shape of hydraulic fractures.

8. The method for coordinated control of hard roof and gas in fully mechanized top caving mining of extra-thick coal seams according to claim 1 is characterized by: The position of the drilling end point is determined according to the gas migration law of the goaf, specifically: The mathematical model of seepage diffusion of gas migration in goaf uses the diffusion equation, gas outflow sink and basic differential equation of seepage as reference functions. According to Bernoulli equation and Darcy theorem, the orifice flow rate of the final hole of the drilling is: q is the orifice flow rate, in m / s; μA is the resistance coefficient of the goaf nozzle; g is the gravity acceleration, in m / s 2 ; z A 、z B are the height of the goaf and the height of the hole respectively; p A is the pressure of the goaf, p B is the pressure at the orifice, λ is g(ρ A -ρ B ), where ρ A is the gas concentration in the goaf, ρ B is the gas concentration at the orifice; k is the permeability coefficient; h is the height, in meters, and s is the length of the seepage path; The diffusion movement of coal seam gas in the goaf follows the diffusion equation in the three-dimensional porous medium, that is, the gas outflow differential equation group, which is: c is the volume percentage concentration of coal seam gas in the goaf; t is time, the unit is s; is the Hamiltonian operator; is the diffusion coefficient; is the wind flow seepage velocity in the goaf; c' is the gas concentration released by the falling gangue; W is the gas distribution source intensity; Substituting the outflow source into the basic differential equations of seepage, we get: Among them, a is ν is the kinematic viscosity coefficient of the mixed gas, the unit is m 2 / s, g is the acceleration due to gravity, the unit is m / s 2 ; b is β is the geometric shape factor, n is the porosity, d is the medium equivalent diameter, the unit is m, e is the permeability scalar coefficient; p is the static pressure; z is the height of the goaf, the unit is m; h 静 is the hydrostatic head; B is the dimension correction coefficient; φ is the porosity in the coal rock mass; C is the isothermal adsorption constant; Where e is a constant, b0 is the decay rate, t0 is the initial time, t1 is the decay time; H is the caving height of the goaf; The porosity of the coal rock mass is assumed to be constant over time and pressure, and the differential equations of the orifice flow rate of the final hole of the drilling, the gas outflow sink, and the basic differential equations of the seepage based on the outflow source are combined to obtain the boundary conditions. Based on the numerical solution, the law of the change of gas concentration in the goaf over time is obtained. According to the judgment of the flow state by Reynolds number, under the original permeability, gas forms linear seepage in the coal seam, and the seepage displacement distance is L1: ρ is the fluid density; k is the permeability; Δp is the pressure difference; μ is the gas dynamic viscosity; Re is the Reynolds number; When the gas flow rate is lower than the flow rate during linear seepage, there is a starting pressure gradient λ0 for the gas. When the pressure gradient is greater than the starting pressure gradient, the gas migration in the coal seam is nonlinear seepage, and the seepage displacement distance is L2: The gas extraction radius is the sum of the seepage displacement distance L1 and the seepage displacement distance L2.

9. The method for coordinated control of hard roof and gas in fully mechanized top caving mining of extra-thick coal seams according to claim 8 is characterized in that: The boundary conditions include: The first boundary condition gives the pressure distribution of gas at the extraction wellhead; The second boundary condition gives the velocity distribution of gas at the extraction wellhead; The third boundary condition gives the relationship between the internal and external pressure difference on the boundary and the gas flow rate.

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