Design method and system for narrow-stop coal pillar main roadway based on stress obstruction

Abstract

The invention belongs to the technical field of coal underground mining, and particularly relates to a design method and system for a narrow coal stopping pillar main roadway based on stress obstruction.The method comprises the steps that the critical stable width of a narrow protection coal pillar is determined according to working face geological and mechanical parameters; based on the influence range of the mining-induced stress and the advancing speed of the working face, the opportunity, the position, the horizon and the area of hydraulic fracturing weakening construction are designed, so that a weakening zone for blocking the transmission of the mining-induced stress is constructed; and according to the key layer theory and the roof structure stability requirement, the angle and height parameters of presplitting blasting roof cutting are determined, and the lateral roof fracture form is optimized. Through combination of stress active blocking and roof directional pressure relief, on the premise of ensuring long-term stability of a main roadway, the width of a protective coal pillar is remarkably reduced, the permanent coal pillar loss is reduced, and meanwhile, the stress concentration effect during short-distance coal seam group mining is relieved.

Description

Technical Field

[0001] This invention relates to the field of underground coal mining technology, and in particular to a design method and system for a narrow stop coal pillar roadway based on stress barrier. Background Technology

[0002] In underground longwall mining of coal, the main roadway undertakes the primary tasks of production, transportation, and ventilation, and its long-term stability is a key concern for coal production enterprises. Traditionally, the width of the stop-mining coal pillar is determined primarily based on the influence distance of the advance support pressure. Therefore, the protective coal pillar in the main roadway is typically in the range of 50-120m to reduce the disturbance impact of the advance support pressure. This results in significant permanent coal resource losses and reduced economic benefits for the enterprise. Furthermore, after the mining of the working faces on both sides of the main roadway is completed, a relatively wide residual coal pillar is formed. During the downward mining of nearby coal seams, an abnormal stress concentration zone forms near the residual coal pillar. If the advance direction of the longwall working face is perpendicular to the axis of the coal pillar during the mining of the lower coal seam, the dynamic load and stress concentration effect transmitted from the coal pillar to the lower working face are significant due to the change in the balance state of the roof structure of the overlying residual coal pillar during the coal pillar extraction process. This can easily lead to safety accidents such as the working face support being crushed and severe coal wall spalling. If the advance direction of the longwall face during the mining of the lower coal seam is parallel to the axis of the overlying coal pillar, in order to reduce the disturbance of the mining roadway by the stress concentration effect of the overlying coal pillar on the goaf, the mining roadway is usually arranged below the goaf at a certain horizontal distance from the boundary of the overlying coal pillar goaf. This not only leads to an increase in the loss of permanent coal pillars during the mining of this coal seam, but also causes the width of the coal pillar to increase continuously in a step-like manner during the mining of lower coal seams, further increasing the loss of permanent coal pillars and seriously interfering with the layout and production system design of longwall mining of the lower coal seam during the mining of multiple coal seams in close proximity.

[0003] Therefore, there is an urgent need for a design method that can effectively protect the stability of the main roadway, significantly reduce coal pillar loss, and improve the mining conditions of closely spaced coal seams. Summary of the Invention

[0004] This invention proposes a design method and system for narrow coal pillar roadways based on stress isolation, aiming to solve the technical problem of stress concentration when mining coal seam groups at a distance from the existing technology.

[0005] In a first aspect, embodiments of the present invention provide a design method for a narrow stop-mining pillar roadway based on stress barrier, comprising: S1, to obtain the production geological conditions and rock strata physical and mechanical parameters of the working face; S2, Based on the aforementioned production geological conditions and rock strata physical and mechanical parameters, determine the critical stable width of the narrow protective coal pillar used to protect the main roadway; S3, based on the principles of reducing mining-induced stress disturbance and ensuring the stability of the roof during equipment withdrawal, determines the timing, location, stratum, and construction area for hydraulic fracturing weakening construction; S4. Based on the principle of optimizing the lateral roof structure to reduce the coal pillar load, the cutting angle and cutting height of the pre-splitting blasting roof cutting are determined. S5, based on the parameters determined in S2 to S4, designs the narrow stop coal pillar roadway.

[0006] The technical effect of the design method of narrow coal pillar roadway based on stress barrier disclosed in this invention is that by constructing a stress barrier through hydraulic fracturing and combining it with pre-splitting blasting to optimize the roof structure, the width of the protective coal pillar can be significantly reduced, resource loss can be reduced, and stress concentration problems during the mining of close-range coal seams can be alleviated, while ensuring the long-term stability of the roadway.

[0007] Furthermore, in S2, determining the critical stable width of the narrow protective coal pillar includes: Calculate the average load P borne by the narrow protective coal pillar: ; in, denoted as the average unit weight of the overlying strata, H as the coal seam burial depth, B as the critical stability width, w1 as the width of the main roadway, w2 as the length of the hydraulic support for roof control at the working face, C1 as the stress transmission coefficient of mining, and C2 as the stress concentration coefficient. Calculate the safety factor SF of the coal pillar: ; Where, σ s The strength of the coal pillar is estimated using the empirical formula for coal pillar strength. The critical stability width B is determined based on the following conditions: SF≥1; ; in, To ensure the effective support width of the anchor bolts, x is the width of the central elastic core zone, and x0 is the plastic width of the goaf side.

[0008] Furthermore, in S3, determining the timing of hydraulic fracturing weakening includes: Make the distance L between the fracturing zone and the working face Z satisfy: ; Where L0 is the range of influence of mining stress, K is the influence coefficient considering hydraulic fracturing construction factors and working face advance speed, V is the working face advance speed, and T is the hydraulic fracturing construction cycle.

[0009] Furthermore, in step S3, determining the location of the hydraulic fracturing weakening operation includes: The horizontal distance W between the boundary of the weakened zone and the solid coal seam of the main roadway satisfies: ; Where w1 is the width of the main roadway, B is the critical stability width, w2 is the jacking length of the hydraulic support at the working face, and L p x1 represents the periodic pressure distance of the longwall working face, and x1 represents the width of the plastic zone on the solid coal side of the main roadway.

[0010] Furthermore, in S3, determining the layer for hydraulic fracturing weakening construction includes: Vertical height H of the hydraulically fracturing layer T satisfy: ; Among them, H C H is the minimum vertical distance between the fracturing height and the coal seam roof, and the weak rock strata below the first critical layer are not weakened. R M is the vertical distance between the lower surface of the first critical layer and the roof of the coal seam, and M is the coal seam mining thickness.

[0011] Furthermore, in S3, determining the construction area for hydraulic fracturing weakening includes: Fan-shaped drilling sites are set up in the return airway and transport roadway of the working face, and fan-shaped boreholes are drilled towards the center of the working face to cover the target fracturing layer; if the coverage by the fan-shaped boreholes is insufficient, additional boreholes are drilled from the main roadway towards the working face.

[0012] Furthermore, in S4, determining the depressurization angle θ for the pre-splitting blasting decapitation includes: ; ; ; ; in, θ is the internal friction angle of the key block, q is the load concentration, L is the lateral span after the basic roof fracture, h is the thickness of the basic roof strata, ΔS is the subsidence after the basic roof fracture, M is the coal seam mining thickness, η is the coal extraction rate of the working face, h z For the direct top height, K p Here, T is the direct crushing expansion coefficient, T is the horizontal friction force of the cantilever beam, and R is the shear force experienced during the unstable sliding process.

[0013] Furthermore, in step S4, the process of determining the cutting height includes: ; Where H2 is the vertical distance between the upper surface of the basic top layer and the roof of the coal seam, H3 is the vertical distance between the upper surface of the high-level key stratum and the roof of the coal seam, M is the coal seam mining thickness, η is the coal extraction rate of the working face, and K p This is the direct crushing expansion coefficient.

[0014] Furthermore, the method also includes S6: when the working face equipment is withdrawn, directional pre-splitting blasting is carried out according to the parameters determined in step S4, and the hydraulic supports are withdrawn in sequence so that the roof of the goaf collapses in time.

[0015] Secondly, embodiments of the present invention provide a design system for a narrow stop-mining pillar roadway based on stress barrier, the system being configured to execute the method, the system comprising: The parameter acquisition module is configured to acquire the production geological conditions and rock strata physical and mechanical parameters of the working face; The coal pillar width calculation module is configured to determine the critical stable width of the narrow protective coal pillar used to protect the main roadway based on the production geological conditions and the physical and mechanical parameters of the rock strata. The hydraulic fracturing parameter design module is configured to determine the timing, location, stratigraphic level, and construction area of ​​hydraulic fracturing weakening operations based on the principles of reducing mining-induced stress disturbance and ensuring the stability of the roof during equipment withdrawal. The pre-splitting blasting parameter design module is configured to determine the cutting angle and cutting height of the pre-splitting blasting roof cutting based on the principle of optimizing the lateral roof structure to reduce the coal pillar load. The process integration module is configured to design the narrow stop coal pillar roadway based on the parameters determined in S2 to S4.

[0016] The technical advantages of the design system for narrow-pillar coal roadways based on stress isolation disclosed in this invention are as follows: The system modularizes and standardizes the complex design process, reducing excessive reliance on human experience and minimizing human error. Through integrated parameter processing and process management, the system ensures the consistency and reliability of design results, improving design efficiency and engineering operability. Its structured design facilitates digital application and decision support in actual mining operations. Attached Figure Description

[0017] Figure 1 A flowchart illustrating a design method for a narrow stop coal pillar roadway based on stress barrier, provided in an embodiment of the present invention; Figure 2 A schematic diagram illustrating the timing of hydraulic fracturing provided in an embodiment of the present invention; Figure 3 This is a schematic diagram of the location of hydraulic fracturing provided in an embodiment of the present invention; Figure 4 This is a schematic diagram of the hydraulic fracturing construction area provided in an embodiment of the present invention; Figure 5 This is a schematic diagram illustrating the determination of pre-splitting blasting cutting parameters provided in an embodiment of the present invention; Figure 6 A schematic diagram illustrating the beneficial technical effects of the narrow stop coal pillar provided in this embodiment of the invention; Figure 7 This is a schematic diagram of a rock stratum column provided for an embodiment of the present invention. Detailed Implementation

[0018] To enable those skilled in the art to better understand the technical solutions of the present invention, exemplary embodiments of the present invention are described below in conjunction with the accompanying drawings, including various details of the embodiments of the present invention to aid understanding. These should be considered merely exemplary. Therefore, those skilled in the art should recognize that various changes and modifications can be made to the embodiments described herein without departing from the scope and spirit of the present invention. Similarly, for clarity and brevity, descriptions of well-known functions and structures are omitted in the following description.

[0019] A design method for narrow stop coal pillar roadways based on stress barrier, referencing Figures 1 to 7 As shown, the specific steps include: S1. Obtain the production geological conditions and physical and mechanical parameters of the rock strata at the working face. Through on-site investigation and detailed coal and rock strata columnar section mapping, and by core drilling and standard laboratory testing, obtain the physical and mechanical parameters of the coal seam and its roof and floor strata, such as uniaxial compressive strength, cohesion, internal friction angle, and unit weight. Simultaneously, through on-site monitoring or theoretical experience, determine key production parameters such as the range of influence of mining-induced stress, the periodic pressure step distance, and the planned advance speed of the working face.

[0020] S2, based on the aforementioned production geological conditions and rock strata physical and mechanical parameters, determine the critical stable width of the narrow protective coal pillar used to protect the main roadway.

[0021] The determination of the critical stability width mainly consists of two parts: assessing the strength of the coal pillar and calculating the bearing load. Ultimately, a safety factor is used to reflect the stability of the protective coal pillar. Ignoring the influence of the rock strata fracture angle, this invention uses the stress transmission coefficient C1 and stress concentration coefficient C2 to characterize the average load P borne by the protective coal pillar. ; in, denoted as the average unit weight of the overlying strata, H as the coal seam burial depth, B as the critical stability width, w1 as the width of the main roadway, w2 as the length of the hydraulic support for roof control at the working face, C1 as the stress transmission coefficient of mining, and C2 as the stress concentration coefficient. Calculate the safety factor SF of the coal pillar: ; Where, σs The strength of the coal pillar is estimated using the empirical formula for coal pillar strength. The critical stability width B is determined based on the following conditions: SF≥1; ; in, To ensure the effective support width of the anchor bolts, x is the width of the central elastic core zone, and x0 is the plastic width of the goaf side.

[0022] This invention uses various empirical formulas for coal pillar strength to verify the reasonable width of narrow coal pillars, as shown in Table 1.

[0023] Table 1 Empirical Formula for Coal Pillar Strength

[0024] Where, σ p This refers to the uniaxial compressive strength of the coal mass as determined in the laboratory.

[0025] The minimum safe width of the narrow coal pillar was scientifically determined from both mechanical load-bearing and engineering practicality perspectives, minimizing coal resource loss while ensuring the protection function of the main roadway.

[0026] S3, based on the principles of reducing the degree of mining-induced stress disturbance during the working face mining period and ensuring the stability of the roof structure during equipment withdrawal after the working face is shut down, determines the timing of hydraulic fracturing, the location of the weakened zone, and the strata of the weakened rock layer. This step aims to create a weakened zone in the rock strata through hydraulic fracturing technology to prevent the transmission of mining-induced stress to the main roadway and to create conditions for subsequent roof cutting and pressure relief.

[0027] (1) Timing of hydraulic fracturing. Hydraulic fractures always extend perpendicular to the direction of the minimum principal stress. In coal-bearing strata with a depth of less than 1000 m, the geostress is mostly dominated by the horizontal principal stress. Therefore, in the early stage of longwall mining (outside the range of mining-induced stress), hydraulic fracturing can create a weakened zone to promote the horizontal extension of hydraulic fractures and weaken the rock mass to a greater extent. Assuming the range of mining-induced stress is L0, the distance L between the fracturing zone and the working face should be L when hydraulic fracturing is implemented to weaken the rock mass. Z Must meet: ; Where K is the influence coefficient considering hydraulic fracturing construction factors and working face advance speed, V is the working face advance speed, T is the hydraulic fracturing construction cycle, and the timing of hydraulic fracturing is as follows: Figure 2 As shown.

[0028] (2) Location of hydraulic fracturing. The determination of the location of hydraulic fracturing should follow the following principles: 1) To minimize the impact of mining stress on the main roadway during the entire mining process, the location of the hydraulic fracturing weakened zone should be as close as possible to the main roadway. 2) To reduce the risk of support crushing and equipment damage due to unreasonable fracture of the basic roof during equipment retreat when the working face advances to the stop mining position. Therefore, to reduce the workload and improve construction efficiency, the hydraulic fracturing roof cutting locations for both the hydraulic fracturing weakened zone construction and the roof structure optimization during the final mining period can be normalized. The location of hydraulic fracturing is as follows: Figure 3 As shown.

[0029] The horizontal distance W between the boundary of the weakened zone and the solid coal seam of the main roadway satisfies: ; Where w1 is the width of the main roadway, B is the critical stability width, w2 is the jacking length of the hydraulic support at the working face, and L p x1 represents the periodic pressure distance of the longwall working face, and x1 represents the width of the plastic zone on the solid coal side of the main roadway.

[0030] (3) Hydraulic fracturing layer.

[0031] The hydraulic fracturing level is determined based on the critical layer theory, and currently the highest fracturing level must be no less than 8 times the coal seam mining thickness. Meanwhile, to avoid adverse effects on the hydraulic supports of the working face due to weakened roof after fracturing during the advancement of the working face into the main roadway, the weak rock strata below the first critical layer (basic roof) are not weakened. Therefore, the hydraulic fracturing level should meet the following requirements: ; Among them, H C H is the minimum vertical distance between the fracturing height and the coal seam roof, and the weak rock strata below the first critical layer are not weakened. R M is the vertical distance between the lower surface of the first critical layer and the roof of the coal seam, and M is the coal seam mining thickness.

[0032] (4) Fracturing Operation Area. To improve construction efficiency, two fan-shaped drilling sites are arranged in the return airway and transport roadway of the working face to fully cover the target fracturing layer within the length of the working face. For the target fracturing layer along the length of the working face, if the fan-shaped boreholes cannot fully cover it, additional boreholes can be drilled in the main roadway to supplement the coverage. See [link to hydraulic fracturing operation area] for details. Figure 4 As shown.

[0033] By precisely designed hydraulic fracturing, a continuous "stress barrier" is formed in the surrounding rock in advance, which actively guides and attenuates mining stress, significantly reducing the static and dynamic loads transmitted to the main roadway, laying the foundation for the stability of the main roadway and the application of narrow coal pillars.

[0034] S4. Based on the principle of optimizing the lateral roof structure to reduce the damage caused by coal pillar load, the pre-splitting blasting roof cutting parameters are determined. Pre-splitting blasting roof cutting and pressure relief mainly includes the cutting angle and cutting height. This step addresses the roof control problem during the working face shutdown and equipment retraction phases. Through directional blasting, the lateral overhang is actively cut off, causing it to collapse in a predetermined manner after equipment retraction.

[0035] To prevent the lateral roof from fracturing at an unreasonable location under overlying loads after the working face equipment is withdrawn, directional pre-splitting blasting technology is used to cut and relieve pressure on the roof during the withdrawal of the hydraulic supports. After the working face supports are withdrawn, the roof collapses in a timely manner, ensuring that the narrow stop coal pillar and main roadway are under the intact cantilever beam roof structure. Figure 5 As shown.

[0036] (1) Cutting angle: The block E must satisfy the following conditions to prevent it from becoming unstable and sliding down after the top is cut by blasting: ; ; ; ; in, θ is the internal friction angle of the key block, q is the load concentration, L is the lateral span after the basic roof fracture, h is the thickness of the basic roof strata, ΔS is the subsidence after the basic roof fracture, M is the coal seam mining thickness, η is the coal extraction rate of the working face, h z For the direct top height, K p Here, T is the direct crushing expansion coefficient, T is the horizontal friction force of the cantilever beam, and R is the shear force experienced during the unstable sliding process.

[0037] (2) Cutting height: The cutting height is related to the coal seam thickness, the position of the hard rock strata, the rock strata fragmentation coefficient, and the coal seam extraction rate, and it must ensure that the basic roof and the high-level key strata can collapse in a timely manner. Therefore, the cutting height must meet the following requirements: ; Where H2 is the vertical distance between the upper surface of the basic top layer and the roof of the coal seam, H3 is the vertical distance between the upper surface of the high-level key stratum and the roof of the coal seam, M is the coal seam mining thickness, η is the coal extraction rate of the working face, and K p This is the direct crushing expansion coefficient.

[0038] The proactive pre-splitting blasting roof cutting transforms the potentially dangerous "inverted triangle" roof structure that might form after equipment retraction into a stable "cantilever beam" structure, and allows the roof to collapse in time to fill the goaf, completely eliminating the potential load threat from the lateral roof to the narrow protective coal pillars and main roadways, and ensuring the long-term low-stress stability of the area.

[0039] S5. Substitute the parameters determined in S1 into the above formula, execute steps S2-S4, determine the critical width B of the coal pillar stability, the parameters of the hydraulic fracturing weakened zone, and the pre-splitting blasting cutting angle and height, and finally determine the above process flow.

[0040] S6, during the withdrawal of the working face equipment, directional pre-splitting blasting is carried out according to the parameters determined in step S4, and the hydraulic supports are withdrawn sequentially to ensure the timely collapse of the goaf roof. Specifically, blasting boreholes are drilled and charged from the roadway side towards the goaf. Group blasting is typically used, and the supports in that area are immediately withdrawn after blasting to ensure the roof collapses in the designed direction. This step is repeated until all supports have been safely withdrawn.

[0041] The relevant technical principles of this invention are as follows: (1) The working principle of the stress weakening zone: Hydraulic fracturing technology is used to weaken rock strata and create artificial faults, disrupting the integrity of the rock layers and forming cavities and fracture zones within them, similar to the effect of a non-drop fault or a collapse column structure. Because the coal and rock mass in the fault fracture zone is relatively fragmented, it has a stress-relieving and pressure-blocking effect. Stress unloading at the fracture zone causes a redistribution of the pre-support stress at the working face, thereby reducing the peak support stress and its impact range.

[0042] Furthermore, during the advancement of the longwall working face, the overlying hard rock strata undergo periodic fracturing, and the dynamic load generated by this fracturing is transmitted downwards in the form of stress waves. Since stress waves attenuate less in hard rock strata but significantly in weak rock strata, the weakened zone can significantly reduce the intensity and energy of the dynamic load transmitted to the main working face roadway.

[0043] (2) Pre-splitting blasting and top cutting pressure relief mechanism: When the longwall face advances to the stop line, the main roof will be in a cantilever beam state and generally will not experience periodic pressure, ensuring the safe withdrawal of equipment. After the equipment withdrawal is completed, the weak direct roof collapses into the goaf, while the lateral main roofs, in an inverted triangle cantilever state, bend and sink under their own weight and mining stress. When the internal stress reaches its ultimate strength, it breaks and rotates downwards. When the main roof fractures above the goaf, the resulting blocks mainly spiral down towards the goaf. The roof structure above the main roadway and narrow coal pillars remains relatively intact (cantilever beam structure), and the additional load from the overlying strata is limited, resulting only in slight roof convergence and shallow yielding of the coal pillars, which is most beneficial to the stability of the main roadway.

[0044] Beneficial technical effects of the present invention: Compared with setting up wide coal pillars to protect the main roadway, the present invention has the following significant advantages: (1) Hydraulic fracturing is used to weaken the target rock strata. The weakened zone can effectively block the transmission of mining stress, which can better reduce the deformation of the main roadway and reduce its repair frequency and workload. (2) Using narrow protective coal pillars can greatly reduce the permanent loss of coal pillar resources during the mining of this coal seam, increase the extraction rate of the mining area and extend the service life of the mine; (3) Using pre-splitting blasting technology to cut and depressurize the lateral roof can effectively avoid the adverse effects of unreasonable fracture of the basic roof on the surrounding rock of the roadway and the narrow protective coal pillar, and ensure the normal service of the main roadway. (4) Narrow protective coal pillars gradually yield and fail after the mining of the working faces on both sides of the main roadway is completed. This can reduce the stress concentration effect caused by the left coal pillars and effectively solve the risk of the support being crushed during the coal pillar production of the lower working face during the mining of close-distance coal seams.

[0045] (5) In addition, after the narrow coal pillar yields and fails, there are more options for the layout of the longwall mining roadway of the lower coal seam. This effectively avoids the problem that the stress concentration of the coal pillar will further increase the protective coal pillar between the longwall mining roadways of the lower working face. It can realize the decompression mining of coal seam groups and optimize the mining layout of the longwall working face during multi-coal seam mining.

[0046] The invention will be further illustrated below with specific application examples: Step S1: The average burial depth of coal seam No. 2 in a certain mine is H=272m, the working face length is 150m, the coal seam mining thickness is M=3.1m, and the fully mechanized one-pass full-height mining method is adopted. The hydraulic support roof control length is w2=8.0m, and the average unit weight of the overlying strata is γ=24.0kN / m³. 3 The main roadway has a cross-sectional width w1 = 4.5m. Anchor-mesh-cable combined support is used for tunneling. The coal and rock strata columnar section is shown below. Figure 7 As shown in the figure. On-site monitoring revealed that the disturbance stress range of the mining stress was 45m, and the working face advance speed was 5m / day.

[0047] Laboratory and field measurements of uniaxial compressive strength σ of coal p =10.8MPa Coal seam internal friction angle φ0=22°, cohesion c0=1.8 MPa, coal seam lateral pressure coefficient λ=0.7, stress concentration factor k1=2.9, support strength p0=0.1MPa.

[0048] Step S2: Determine the critical width for maintaining a stable coal pillar to ensure that narrow coal pillars remain stable during equipment retraction after the working face is shut down. Based on the above formula, Table 1, and the geological conditions of production, the parameters are selected as follows: —Average unit weight of overlying strata, 24 kN / m³ 3 ; H—Deep burial depth of the coal seam, 272m; w1—Main lane width, 4.5m; w2—Length of hydraulic support for jacking at the working face, 8.0m; C1—Transmission coefficient of mining stress, 0.2; C2—Stress concentration factor, 3.5 M—Height of the coal pillar, 3.1m.

[0049] Using the coal pillar safety factor SF=1 as the criterion, the critical width of the coal pillar under different coal pillar strength formulas is: (A1)B=7.48m, (A2)B=5.75m, (A3)B=5.23m.

[0050] In addition, based on engineering experience, the lower limit W of the width of a narrow protective coal pillar can maintain stability and isolate the goaf. p The above formula should be satisfied. The parameters are selected as follows: W p1 —Effective support width of the side anchor bolts: 2.0m; K s —This represents the coefficient for the central elastic core region, 0.2; λ—lateral pressure coefficient, 0.7; c0—cohesion within the coal seam, 1.8 MPa; φ0—Internal friction angle of the coal seam, 22°; k1—The concentration factor of the supporting stress in the coal seam, 2.9; p0—Strength of side anchor bolt support in the goaf of narrow protective coal pillar, 0.1 MPa; H—Deep burial depth of the coal seam, 272m; Substituting the above parameters into the above formula, the lower limit W of the critical stability width is obtained through theoretical calculation. p It is 7.64 m.

[0051] Based on the above analysis and according to the above formula, the width B of the narrow coal pillar is determined to be 8.0m.

[0052] Note: This invention considers the additional load from lateral roof overhang when calculating the load borne by the narrow coal pillar (the hydraulic support roof control length w2 = 8.0m, forming a roof overhang after the support is withdrawn). Because pre-splitting blasting is performed to remove pressure from the lateral roof overhang during the withdrawal of the hydraulic support at the working face, the additional load from the lateral roof overhang of the narrow coal pillar is eliminated, resulting in higher safety for the 8.0m wide protective coal pillar.

[0053] Step S3: Based on the principles of reducing the degree of mining-induced stress disturbance during the working face mining period and ensuring the stability of the roof structure during the equipment withdrawal period after the working face is shut down, determine the timing of hydraulic fracturing construction, the location of the weakened zone, and the stratigraphic position of the weakened rock layer.

[0054] (1) Timing of hydraulic fracturing: The stress influence range of this mining operation is L0, which is 45m. The working face advance speed V is 5m / day, and the hydraulic fracturing construction cycle T is 25 days. Considering the influence coefficient K=1.2, substituting the above parameters into the above formula, we can obtain the distance L between the fracturing zone and the working face. Z ≥45+5×25×1.2=135m.

[0055] (2) Location of hydraulic fracturing: The width B of the narrow stop coal pillar is 8.0m, the hydraulic support roof control length w2 is 8.0m, the width x1 of the plastic zone on the solid coal side of the main roadway is 4.1m, and the periodic pressure distance L of the longwall working face is... p The distance is 25.7 m. Therefore, substituting the above parameters into the above formula, we can obtain that the reasonable horizontal distance W between the boundary of the weakened zone and the solid coal bed of the main roadway should satisfy: 4.5 + 8.0 + 8.0 = 20.5 m ≤ W ≤ 25.7 - 4.1 = 21.6 m.

[0056] (3) Hydraulic fracturing layers: This working face mines the No. 2 coal seam. The area between the main roof and the roof of the No. 2 coal seam is a weak rock stratum, located at a distance of H... C The basic top fine sandstone layer is 6.2m thick, the key layer is 8.2m thick, and the vertical distance between the basic top and the key layer is 9.6m (HR=30.5m). The coal seam mining thickness is 3.1m. The eight-fold mining height is 24.8m. Therefore, according to the above formula, the maximum hydraulic fracturing height HR is 30.5m, and the minimum height H is... C The vertical distance is 6.2m. Therefore, the target fracturing layer is the rock stratum with a vertical height of 6.2m to 30.5m from the top of the coal seam.

[0057] (4) Fracturing operation area: Two fan-shaped drilling sites were set up in the return airway and transport roadway of the working face to fully cover the target fracturing layer within the length of the working face. The distance from the left boundary of the fracturing area to the main roadway return slope is (B+w2)16m.

[0058] Step S4: Based on the principle of optimizing the lateral roof structure to reduce the damage caused by coal pillar load, determine the pre-splitting blasting roof cutting parameters.

[0059] The basic roof thickness h is 8.2 m, H2 is 14.4 m, and H3 is 30.5 m. Based on the weakening position (W) of hydraulic fracturing, 8.0 m ≤ L ≤ 8.0 + 21.6 - 20.5 = 9.1 m. The coal recovery rate η of the working face is taken as 93%, and the immediate roof height h... z The diameter is 6.2 m, and the direct roof expansion coefficient Kp is taken as 1.2. φ1 is taken as 45°. Based on the above parameters and the above formula, θ and H can be obtained. zThey are respectively: θ≥9.33°~12.75°, H z ≥30.5 m.

[0060] The following is a reiteration of the specific technological process of the design method for protecting narrow coal pillars in roadways based on stress isolation and roof cutting pressure relief: (1) The width of the narrow stop-mining protective coal pillar is designed to be 8.0m. When the working face advances to a point where the horizontal distance between it and the main roadway is not less than L, Z When +w2+B=135+8+8=151m, a fan-shaped drilling site is set up in the return airway and transport roadway at a reasonable horizontal distance W from the solid coal seam in the main roadway, respectively, to carry out hydraulic fracturing. Among these, 4.5+8.0+8.0=20.5m≤W≤25.7-4.1=21.6 m, the highest fracturing layer is 30.5 m, and the minimum fracturing layer is 6.2 m from the coal seam roof.

[0061] (2) When the working face is advanced to the stop line position (the horizontal distance between the working face and the main roadway is 8.0m), the mining equipment is withdrawn. After the coal mining machine, scraper conveyor and other equipment are withdrawn, with θ as 10° and the maximum cutting height of the top is 30.5m, the blasting boreholes are constructed sequentially from the side of the return airway, with a borehole spacing of 1000mm.

[0062] (3) After drilling is completed, blasting work is carried out sequentially starting from the side of the transport roadway. Each blast consists of three boreholes as a group, numbered 1, 2, and 3. Borehole No. 2 is a pilot hole (located between boreholes No. 1 and No. 3) and is not loaded with explosives. Boreholes No. 1 and No. 3 are detonated simultaneously. After blasting, the hydraulic supports in the blasting end area are sequentially withdrawn, and the roof of the goaf collapses in time to reduce the load on the narrow protective coal pillar. The blasting of all boreholes and the withdrawal of hydraulic supports are completed in this order.

[0063] The following example illustrates the specific parameters of this invention during implementation, using a particular scenario as an example. It should be noted that the technical solution of this invention is not limited to the specific application examples below: The hydraulic fracturing boreholes and pre-fracturing blasting boreholes have a diameter of 65 mm, with a maximum vertical height of 30.5 m at the borehole termination point. The spacing between hydraulic fracturing points within a single borehole is between 3.0 m and 3.5 m. The sealing device is a double-ended capsule seal type, model HF-120 MPa, with a diameter of 55 mm. The water injection equipment uses a high-pressure plunger pump with a flow rate of 120 L / min and a maximum pressure of 70 MPa. Considering the working face length of 120 m, and based on symmetry, the hydraulic fracturing boreholes H1 to H7 are arranged consistently in the transport roadway and return airway, with drilling angles of 65°, 55°, 45°, 35°, 28°, 20°, and 15°, and lengths of 33.5 m, 37.0 m, 43.0 m, 53.5 m, 66.0 m, 62.0 m, and 59.5 m, respectively.

[0064] The pre-splitting blasting uses BTC-1500 shaped charge tubes with an outer diameter of 42mm, an inner diameter of 36.5mm, and a single shaped charge tube length of 1.5m. Class III coal mine emulsion explosives are used, with cartridge specifications of φ35×300mm, 300g / cartridge, and a charge load of 1.5kg / m in the shaped charge tube. φ50mm×200mm water-based drilling mud is used for hole sealing, with a sealing length not less than 1 / 3 of the borehole length. The total borehole length for pre-splitting blasting is 31m, with a borehole spacing of 2.0m, for a total of 61 boreholes. The pre-splitting blasting process flow is as follows: Safety monitoring → Water spraying → Charging and connection → Checking gas and CO2 concentrations → Water spraying → Blasting → Safety monitoring → Water spraying → Safety monitoring → Next cycle.

[0065] Based on the same inventive concept, embodiments of the present invention also provide a design system for a narrow stop-mining pillar roadway based on stress barrier, the system being configured to execute the method, the system comprising: The parameter acquisition module is configured to acquire the production geological conditions and rock strata physical and mechanical parameters of the working face; The coal pillar width calculation module is configured to determine the critical stable width of the narrow protective coal pillar used to protect the main roadway based on the production geological conditions and the physical and mechanical parameters of the rock strata. The hydraulic fracturing parameter design module is configured to determine the timing, location, stratigraphic level, and construction area of ​​hydraulic fracturing weakening operations based on the principles of reducing mining-induced stress disturbance and ensuring the stability of the roof during equipment withdrawal. The pre-splitting blasting parameter design module is configured to determine the cutting angle and cutting height of the pre-splitting blasting roof cutting based on the principle of optimizing the lateral roof structure to reduce the coal pillar load. The process integration module is configured to design the narrow stop coal pillar roadway based on the parameters determined in S2 to S4.

[0066] The preferred embodiments of the present invention have been described in detail above. However, the present invention is not limited to the specific details of the above embodiments. Within the scope of the technical concept of the present invention, various equivalent transformations can be made to the technical solution of the present invention, and all such equivalent transformations fall within the protection scope of the present invention. The present invention focuses on the specific process of implementing the method and the determination of key parameters. Detailed descriptions are not provided for hydraulic fracturing drilling, water injection time, water injection fracturing, borehole sealing device type, blasting top cutting borehole diameter, and charge quantity. The hydraulic fracturing technology used in the present invention is conventional hydraulic fracturing technology, and the corresponding blasting parameters can be adjusted according to the actual conditions of the mine.

[0067] Example embodiments have been disclosed herein, and while specific terminology has been used, it is for illustrative purposes only and should be construed as such, and is not intended to be limiting. In some instances, it will be apparent to those skilled in the art that features, characteristics, and / or elements described in conjunction with particular embodiments may be used alone, or in combination with features, characteristics, and / or elements described in conjunction with other embodiments, unless otherwise expressly indicated. Therefore, those skilled in the art will understand that various changes in form and detail may be made without departing from the scope of the invention as set forth in the appended claims.

Claims

1. A design method for narrow-stop coal pillar roadways based on stress barrier, characterized in that, include: S1, obtain the production geological conditions and rock strata physical and mechanical parameters of the working face; S2, Based on the aforementioned production geological conditions and rock strata physical and mechanical parameters, determine the critical stable width of the narrow protective coal pillar used to protect the main roadway; S3, based on the principles of reducing mining-induced stress disturbance and ensuring the stability of the roof during equipment withdrawal, determines the timing, location, stratum, and construction area for hydraulic fracturing weakening construction; S4. Based on the principle of optimizing the lateral roof structure to reduce the coal pillar load, the cutting angle and cutting height of the pre-splitting blasting roof cutting are determined. S5, based on the parameters determined in S2 to S4, designs the narrow stop coal pillar roadway.

2. The method according to claim 1, characterized in that, In S2, determining the critical stable width of the narrow protective coal pillar includes: Calculate the average load P borne by the narrow protective coal pillar: ； in, denoted as the average unit weight of the overlying strata, H as the coal seam burial depth, B as the critical stability width, w1 as the width of the main roadway, w2 as the length of the hydraulic support for roof control at the working face, C1 as the stress transmission coefficient of mining, and C2 as the stress concentration coefficient. Calculate the safety factor SF of the coal pillar: ； Where, σ s The strength of the coal pillar is estimated using the empirical formula for coal pillar strength. The critical stability width B is determined based on the following conditions: SF≥1; ； in, To ensure the effective support width of the anchor bolts, x is the width of the central elastic core zone, and x0 is the plastic width of the goaf side.

3. The method according to claim 1, characterized in that, In step S3, determining the timing of hydraulic fracturing weakening includes: Make the distance L between the fracturing zone and the working face Z satisfy: ; Where L0 is the range of influence of mining stress, K is the influence coefficient considering hydraulic fracturing construction factors and working face advance speed, V is the working face advance speed, and T is the hydraulic fracturing construction cycle.

4. The method according to claim 1, characterized in that, In step S3, determining the location for hydraulic fracturing weakening operations includes: The horizontal distance W between the boundary of the weakened zone and the solid coal seam of the main roadway satisfies: ； Where w1 is the width of the main roadway, B is the critical stability width, w2 is the jacking length of the hydraulic support at the working face, and L p x1 represents the periodic pressure distance of the longwall working face, and x1 represents the width of the plastic zone on the solid coal side of the main roadway.

5. The method according to claim 1, characterized in that, In step S3, determining the layer for hydraulic fracturing weakening construction includes: Vertical height H of the hydraulically fracturing layer T satisfy: ； Among them, H C H is the minimum vertical distance between the fracturing height and the coal seam roof, and the weak rock strata below the first critical layer are not weakened. R M is the vertical distance between the lower surface of the first critical layer and the roof of the coal seam, and M is the coal seam mining thickness.

6. The method according to claim 1, characterized in that, In step S3, determining the construction area for hydraulic fracturing weakening includes: Fan-shaped drilling sites are set up in the return airway and transport roadway of the working face, and fan-shaped boreholes are drilled towards the center of the working face to cover the target fracturing layer; if the coverage by the fan-shaped boreholes is insufficient, additional boreholes are drilled from the main roadway towards the working face.

7. The method according to claim 1, characterized in that, In step S4, determining the decompression angle θ for pre-splitting blasting includes: ； ； ； ； in, θ is the internal friction angle of the key block, q is the load concentration, L is the lateral span after the basic roof fracture, h is the thickness of the basic roof strata, ΔS is the subsidence after the basic roof fracture, M is the coal seam mining thickness, η is the coal extraction rate of the working face, h z For the direct top height, K p Here, T is the direct crushing expansion coefficient, T is the horizontal friction force of the cantilever beam, and R is the shear force experienced during the unstable sliding process.

8. The method according to claim 1, characterized in that, In step S4, the process of determining the cutting height includes: ； Where H2 is the vertical distance between the upper surface of the basic top layer and the roof of the coal seam, H3 is the vertical distance between the upper surface of the high-level key stratum and the roof of the coal seam, M is the coal seam mining thickness, η is the coal extraction rate of the working face, and K p This is the direct crushing expansion coefficient.

9. The method according to claim 1, characterized in that, The method further includes S6: when the equipment at the working face is withdrawn, directional pre-splitting blasting is carried out according to the parameters determined in step S4, and the hydraulic supports are withdrawn in sequence so that the roof of the goaf collapses in time.

10. A design system for narrow-stop coal pillar roadways based on stress barrier, characterized in that, The system is configured to perform the method as described in any one of claims 1 to 9, the system comprising: The parameter acquisition module is configured to acquire the production geological conditions and rock strata physical and mechanical parameters of the working face; The coal pillar width calculation module is configured to determine the critical stable width of the narrow protective coal pillar used to protect the main roadway based on the production geological conditions and the physical and mechanical parameters of the rock strata. The hydraulic fracturing parameter design module is configured to determine the timing, location, stratigraphic level, and construction area of ​​hydraulic fracturing weakening operations based on the principles of reducing mining-induced stress disturbance and ensuring the stability of the roof during equipment withdrawal. The pre-splitting blasting parameter design module is configured to determine the cutting angle and cutting height of the pre-splitting blasting roof cutting based on the principle of optimizing the lateral roof structure to reduce the coal pillar load. The process integration module is configured to design the narrow stop coal pillar roadway based on the parameters determined in S2 to S4.

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