Deep hole blasting roof breaking pressure relief method for lateral stress regulation and control of hard rock mine goaf

Abstract

The invention relates to the technical field of mining, in particular to a deep hole blasting roof breaking pressure relief method for lateral stress regulation and control of a hard rock mine goaf, which comprises the following steps: S100, obtaining a vertical distance H from a key layer to an ore body; s200, determining the position h1 of the advance bearing pressure peak value of the working face; the advance bearing pressure influence range is h2; the lateral bearing pressure peak value position of the goaf is h3; the lateral bearing pressure influence range of the goaf is h4; s300, the position of blasting pressure relief is determined in the key layer; s400, a plurality of drill holes are drilled in the main control layer according to the positions in the step S300, the drill holes are filled with explosives, and after hole sealing, the key layer is cut off through a blasting method; and S500, the stress and energy evolution law of the working face after the top plate is blasted are monitored in real time, and then the designed drilling hole layout distance is optimized. On the premise of keeping the completeness of roadway surrounding rock, the transmission path of the overlying rock load of the goaf towards the working face is blocked, so that the advanced mining-induced stress influence of the working face is weakened.

Description

Technical Field

[0001] The invention relates to the technical field of mining, in particular to a deep hole blasting roof breaking and pressure relief method for regulating lateral stress in a goaf of a hard rock mine. Background Art

[0002] Mining is the process of extracting ore raw materials from underground or on the surface. It is an important and complex process in the mining industry, involving many aspects of technology and must be carried out under the premise of ensuring safety and environmental protection. The mining depth of mines in my country has gradually increased, resulting in increased ground stress in mines. This has caused more and more working faces to face stress concentration problems and is prone to ground pressure accidents.

[0003] The pressure relief methods in the existing technology mainly include hydraulic fracturing pressure relief, conventional blasting pressure relief, roof pre-splitting pressure relief and other methods. If hydraulic fracturing is used for pressure relief, it will lead to poor compatibility between the fracturing fluid and the formation, large filtration loss in dense rock formations (such as muddy overburden), and difficulty in effective penetration of the fracturing fluid, resulting in excessively high wellhead pressure but failure to open the formation. It will also lead to attenuation of the proppant conductivity and damage to the natural fracture system. If conventional blasting is used for pressure relief, it will lead to incomplete pressure relief, repeated hole filling is required, delaying the mining progress, and leading to complex processes, high safety risks, and affecting the continuity of mining. If the roof pre-splitting pressure relief method is used, it is difficult to break the thick and hard roof. The increase in the cantilever length will significantly amplify the wall support pressure, exceeding the strength limit of the filling body, resulting in wall collapse and doubling of the tunnel deformation. It will also cause the pre-splitting zone and the roof fracture surface to be misaligned, making it difficult to form a continuous weakening surface.

[0004] Mining at the working face causes advanced abutment pressure in the overlying strata of the goaf. This is especially true when working on one side of the goaf. The lateral abutment pressure in the side goaf, combined with the advanced abutment stress of the working face, can cause severe damage to the roadways along the goaf. The distribution of abutment stress is influenced by many factors, such as the depth of the working face, the thickness of the basic roof, the thickness of the immediate roof, and the lithology, and it changes continuously as the working face advances.

[0005] When a roadway is exposed to extended mining pressure, it poses a significant threat to its stability and safety. This "advanced mining pressure" refers to a stress concentration phenomenon caused by the disruption of the mechanical equilibrium of the rock strata within a certain area ahead of the working face during mining. This sustained stress can lead to severe deformation in the roadway, primarily including spalling, roof falls, and floor heave. Spalling refers to the collapse of the rock mass on either side of the roadway due to the inability to withstand the surrounding stress; roof falls refer to the collapse of the rock mass at the top of the roadway; and floor heave refers to the upward bulging of the roadway floor due to pressure. If these deformations are not promptly and effectively controlled, they can easily lead to major accidents such as water inrush, rockbursts, and gas outbursts. For example, roof collapse may destroy the normal structure of the tunnel, causing groundwater to flow into the tunnel and causing water inrush accidents; rock burst is due to the sudden release of high stress accumulated inside the rock formation, causing rock blocks to be suddenly ejected, posing a great threat to personnel and equipment; gas outburst occurs in high-pressure gas areas, and due to stress changes, gas suddenly gushes out in large quantities, which can easily cause serious consequences such as explosions, thereby posing a great risk to the life safety of personnel, and also causing huge losses to the mine's equipment, facilities and resources, affecting the normal production order of the mine, and bringing huge economic burdens and social impacts.

[0006] Therefore, technicians in this field are committed to developing a deep-hole blasting and roof-breaking pressure relief method for lateral stress regulation in hard rock mine goafs, which is beneficial for blocking the transmission path of the overlying rock load in the goaf toward the working face while maintaining the integrity of the tunnel surrounding rock, so as to reduce the impact of the advanced mining stress on the working face, thereby greatly improving the tunnel stability and ensuring the purpose of safe mining. Summary of the Invention

[0007] The technical problem to be solved by the present invention is to provide a deep hole blasting and roof breaking pressure relief method for lateral stress regulation in the goaf of hard rock mines, which is beneficial to blocking the transmission path of the overlying rock load in the goaf to the working face while maintaining the integrity of the tunnel surrounding rock, so as to reduce the influence of the advanced mining stress on the working face, thereby greatly improving the stability of the tunnel and ensuring the purpose of safe mining. The technical solution of the present invention to solve the above technical problems is as follows: A deep hole blasting roof breaking and pressure relief method for lateral stress regulation in a hard rock mine goaf comprises the following steps: S100. Based on the mine geological data or the drill hole histogram obtained from the rock coring experiment, the number of layers, thickness, lithology and physical and mechanical parameters of the rock layer above the working face ore body are obtained. Based on the previous observation and analysis of ground settlement and roof collapse in the goaf area, the vertical distance H between the key layer and the ore body is obtained; S200. Based on the overburden parameters obtained from the drilling histogram, a numerical simulation model of the working face and the goaf on one side is established, and the microseismic monitoring data is combined to determine Peak position of leading support pressure of working face h1; Leading support pressure influence range h2; Peak position of lateral support pressure in goaf h3; The influence range of lateral support pressure in goaf h4; S300. Determine the location of blasting decompression in the key layer; S400. According to the position in step S300, multiple boreholes are drilled into the main control layer, and explosives are filled in the boreholes. After sealing the holes, blasting is used to cut off the key layer. S500. Real-time monitoring of the stress and energy evolution of the working face after roof blasting, and then optimizing the design of drilling layout distance.

[0008] The beneficial effects of adopting the above scheme are: based on the mine geological data or the drill hole histogram obtained from the rock core sampling experiment, the number of layers, thickness, lithology and physical and mechanical parameters of the rock layer above the working face ore body are fully understood, and the vertical distance H of the key layer from the ore body is determined in combination with the relevant observation data of the previous goaf area. On this basis, a numerical simulation model is established using the overburden parameters and combined with the microseismic monitoring data to accurately determine the peak position h1 and the impact range h2 of the leading support pressure of the working face, as well as the peak position h3 and the impact range h4 of the lateral support pressure of the goaf. This provides a scientific and accurate basis for determining the subsequent blasting pressure relief position, ensuring that the blasting operation can act on the key area in a targeted manner and effectively improving the pressure relief effect; By determining the blasting pressure relief position in the key layer and drilling multiple boreholes into the main control layer, blasting after filling the boreholes with explosives, the key layer can be cut off, blocking the transmission path of the overlying rock load in the goaf to the working face, thereby significantly reducing the impact of the working face on the stress caused by the advanced mining of the working face, and effectively avoiding problems such as severe deformation of the tunnel caused by support stress concentration, such as spalling, roof collapse and floor heave, greatly reducing the risk of major accidents such as water gushing, rock burst and gas outburst, and effectively ensuring the stability of the tunnel and the safety of the mining process; Real-time monitoring of the stress and energy evolution of the working face after roof blasting is carried out, and the drilling layout distance is timely optimized based on the monitoring data, so that the blasting pressure relief plan can be dynamically adjusted according to the actual stress changes, further improving the pressure relief effect. At the same time, it can also avoid problems such as resource waste or insufficient pressure relief caused by unreasonable drilling layout, ensure the long-term stability and safety of the tunnel during the mining process, and facilitate safe and efficient mining operations.

[0009] On the basis of the above technical solution, the present invention can also be improved as follows.

[0010] Furthermore, step S300 specifically includes the following steps: S301. In front of the cut, the working face is of length L, and blasting holes are drilled every 10m, where: Number of drilling holes: N1=L / 10; Drilling angle: α1=arctanα6; α6=H / h1; Drilling depth: l1=H / sinα1; S302. Along both sides of the empty roadway in the working face advance support pressure peak position h1 outward arrangement, the arrangement range is open cut to the working face advance support pressure influence range h2; During the excavation period, the tunnel is drilled every 15m along the empty side. During the mining period, one drill hole is drilled every 15m. During the mining period, two drill holes are drilled every 15m on the side of the mining area. The distance between the tunnel along the empty side and the goaf is l, where Airside: Number of drilling holes: N2= h2 / 7.5; Drilling angle: α2=arctanα7; α7=H / l; Drilling depth: l2=H / sinα2; Mining side: Number of drilling holes 1: N3 = h2 / 15 Drilling angle 1: α3=arctanα8; α8=H / h3; Drilling depth 1: l3=H / sinα3; Number of drilling holes 2: N4 = N3; Drilling 2 angle: α4=arctanα9; α9=H / h4; Drilling depth 2: l4=H / sinα4; S303. Solid pillar tunnels are arranged outward from the peak position h1 of the working face's advanced support pressure. The layout range is from the opening of the cut to the working face's advanced support pressure influence range h2. Blasting pressure relief holes are drilled every 15 m along the working face's inclination along the mining sidewalls, using the same layout parameters as those for side drilling holes 2 along the gob-side tunnels. In addition, based on field anti-bumping experience, drill holes are arranged every 7.5 m in the middle of the tunnel along the working face's inclination. Number of drilling holes: N5 = h2 / 7.5; Drilling angle: α5=80°; Drilling depth: l5=H / sinα5.

[0011] The beneficial effects of adopting the above further solution are: determining specific drilling parameters based on the working face length, parameters related to the advance support pressure, etc., achieving accurate arrangement of drilling holes, ensuring that the blasting energy can accurately act on the key rock formation, effectively improving the effect of blasting top fracture pressure relief, and avoiding resource waste or insufficient pressure relief caused by blind arrangement of drilling holes; The arrangement of drilling holes in the solid pillar tunnel is not only based on the advanced support pressure parameters, but also combined with the on-site anti-bumping experience to arrange drilling holes in the middle of the tunnel along the working face. The integration of theoretical calculations and practical experience makes the drilling arrangement more scientific and reasonable, further enhances the pressure relief effect, and improves the stability and safety of the tunnel.

[0012] Furthermore, in step S400, the borehole diameter is 76 mm to 89 mm, an axial continuous coupled charging method is adopted in the hole, the explosive is an emulsion explosive or a water gel explosive, and the length of the grouting sealing section is ≥ 1 / 3 of the hole depth.

[0013] The beneficial effects of adopting the above further solution are: the appropriate drilling diameter provides specific parameter standards for actual operation, facilitates construction operation and equipment selection, ensures the smooth progress and quality control of drilling operations, and improves the reliability and consistency of blasting operations; Grouting and sealing ensures the sealing quality, prevents blasting gas leakage, improves blasting effect and safety, and enhances the blasting ability to cut off key layers and the pressure relief effect.

[0014] Furthermore, in step S500, the real-time monitoring target is: (a) Drilled stress gauges are placed on the roadway roof to monitor the expansion of the plastic zone. The crushing zone extends to within 5 m from the roadway wall, and the stress peak shifts to a position ≥ 12 m deep in the ore body. (b) Continuous observation using a microseismic monitoring system, requiring an energy release rate of ≤10 3 J; (c) The stent pressure monitoring system records the cyclic pressure step in real time, and the step shortening amplitude is ≥5%.

[0015] The beneficial effect of adopting the above-mentioned further scheme is: combining multiple monitoring means such as borehole stress gauges, microseismic monitoring systems and support pressure monitoring systems, real-time monitoring of key parameters such as stress, microseismicity and support pressure of the working face after blasting can be carried out, which can comprehensively and accurately obtain the evolution law of stress and energy of the working face after roof blasting, and provide timely and reliable data support for the subsequent optimization design of drilling layout distance.

[0016] Further, in step S303: The arrangement of drilling holes for the mining side of the solid pillar roadway must meet the following requirements: The drilling spacing is related to the working face advancement speed v. When v>5m / d, the spacing is adjusted to 10m.

[0017] The beneficial effect of adopting the above-mentioned further scheme is that the dynamic adjustment method can better adapt to the stress changes during the advancement of the working face, ensure that blasting pressure relief can be carried out in a timely and effective manner at different advancement speeds, improve the timeliness and effectiveness of pressure relief, and ensure the stability of the tunnel during rapid advancement.

[0018] Furthermore, after step S500, the following steps are further included: S601. Gob-side tunnels shall be supported by a combination of anchor cables and retractable U-shaped steel, with anchor cable preload ≥ 200 kN. S602. Add φ153mm pressure relief holes in the base plate with an inclination of -45°, a depth of 8m to 10m, and a spacing of 3m to 5m.

[0019] The beneficial effects of adopting the above further scheme are: the support can effectively improve the support strength and stability of the tunnel, enhance the tunnel's ability to resist deformation, provide more reliable support guarantees for the tunnel after blasting and unloading, and reduce the risk of tunnel collapse. BRIEF DESCRIPTION OF THE DRAWINGS

[0020] Figure 1 This is a flowchart of a specific embodiment of the present invention; Figure 2 This is a plan view of the drilling arrangement of a specific embodiment of the present invention; Figure 3 This is a working surface inclined drilling arrangement diagram of a specific embodiment of the present invention; Figure 4 This is a diagram of the arrangement of inclined drilling holes along the empty side of a goaf roadway according to a specific embodiment of the present invention; Figure 5 This is a cross-sectional view of the arrangement of drilling holes along the direction of a solid pillar tunnel according to a specific embodiment of the present invention; Figure 6 This is a cross-sectional view of the drilling arrangement in the direction of the cut hole according to a specific embodiment of the present invention. DETAILED DESCRIPTION

[0021] The principles and features of the present invention are described below with reference to the accompanying drawings. The examples given are only used to explain the present invention and are not used to limit the scope of the present invention.

[0022] In the description of the present invention, it should be understood that the terms "center", "length", "up", "down", "front", "back", "left", "right", "vertical", "horizontal", "top", "inside", "outside", "peripheral", "circumferential" and the like to indicate orientations or positional relationships based on the orientations or positional relationships shown in the accompanying drawings, and are only for the convenience of describing the present invention and simplifying the description, and do not indicate or imply that the referred system or element must have a specific orientation, be constructed and operated in a specific orientation, and therefore should not be understood as a limitation on the present invention.

[0023] In the description of the present invention, “a plurality of” means at least two, for example, two, three, etc., unless otherwise clearly and specifically defined.

[0024] In the present invention, unless otherwise specified or limited, the terms "installed," "connected," "connect," "fixed," etc. should be understood in a broad sense. For example, they can refer to fixed connection, detachable connection, or integration; mechanical connection, electrical connection; direct connection, or indirect connection through an intermediate medium; internal communication between two components, or interaction between two components, unless otherwise specified. Those skilled in the art will understand the specific meanings of the above terms in the present invention based on specific circumstances.

[0025] like Figure 1 、 Figure 2 、 Figure 3 、 Figure 4 、 Figure 5 and Figure 6 As shown, a deep hole blasting roof breaking and pressure relief method for lateral stress control in a goaf of a hard rock mine comprises the following steps: S100. Based on the mine geological data or the drill hole histogram obtained from the rock coring experiment, the number of layers, thickness, lithology and physical and mechanical parameters of the rock layer above the working face ore body are obtained. Based on the previous observation and analysis of ground settlement and roof collapse in the goaf area, the vertical distance H between the key layer and the ore body is obtained; S200. Based on the overburden parameters obtained from the drilling histogram, a numerical simulation model of the working face and the goaf on one side is established, and the microseismic monitoring data is combined to determine Peak position of leading support pressure of working face h1; Leading support pressure influence range h2; Peak position of lateral support pressure in goaf h3; The influence range of lateral support pressure in goaf h4; S300. Determine the location of blasting decompression in the key layer; The specific steps include: S301. In front of the cut, the working face is of length L, and blasting holes are drilled every 10m, where: Number of drilling holes: N1=L / 10; Drilling angle: α1=arctanα6; α6=H / h1; Drilling depth: l1=H / sinα1; S302. Along both sides of the empty roadway in the working face advance support pressure peak position h1 outward arrangement, the arrangement range is open cut to the working face advance support pressure influence range h2; During the excavation period, the tunnel is drilled every 15m along the empty side. During the mining period, one drill hole is drilled every 15m. During the mining period, two drill holes are drilled every 15m on the side of the mining area. The distance between the tunnel along the empty side and the goaf is l, where Airside: Number of drilling holes: N2= h2 / 7.5; Drilling angle: α2=arctanα7; α7=H / l; Drilling depth: l2=H / sinα2; Mining side: Number of drilling holes 1: N3 = h2 / 15 Drilling angle 1: α3=arctanα8; α8=H / h3; Drilling depth 1: l3=H / sinα3; Number of drilling holes 2: N4 = N3; Drilling 2 angle: α4=arctanα9; α9=H / h4; Drilling depth 2: l4=H / sinα4; S303. Solid pillar tunnels are arranged outward from the peak position h1 of the working face's advanced support pressure. The layout range is from the opening of the cut to the working face's advanced support pressure influence range h2. Blasting pressure relief holes are drilled every 15 m along the working face's inclination along the mining sidewalls, using the same layout parameters as those for side drilling holes 2 along the gob-side tunnels. In addition, based on field anti-bumping experience, drill holes are arranged every 7.5 m in the middle of the tunnel along the working face's inclination. Number of drilling holes: N5 = h2 / 7.5; Drilling angle: α5=80°; Drilling depth: l5=H / sinα5.

[0026] In a specific embodiment, based on numerical simulation results, when the borehole spacing is less than 10m, the pressure relief effect is not significantly improved, but the cost increases significantly. When the borehole spacing is greater than 15m, the stress concentration zone cannot be effectively covered, so a spacing of 10m to 15m is preferred. Furthermore, based on field tests and microseismic monitoring data, it was found that a spacing of 7.5m can effectively block the stress transmission path in the critical layer. Combined with the blasting impact radius R (obtained through experiments or empirical formulas), the borehole spacing is determined to be ≤ 2R, with a preferred spacing of 7.5m / 10m / 15m to meet this requirement.

[0027] The arrangement of drilling holes for the mining side of the solid pillar roadway must meet the following requirements: The drilling spacing is related to the working face advancement speed v. When v>5m / d, the spacing is adjusted to 10m.

[0028] S400. According to the position in step S300, multiple boreholes are drilled into the main control layer, and explosives are filled in the boreholes. After sealing the holes, blasting is used to cut off the key layer. In step S400, the borehole diameter is 76 mm to 89 mm, the hole is charged in an axial continuous coupled manner, the explosive is an emulsion explosive or a water gel explosive, and the length of the grouting sealing section is ≥ 1 / 3 of the hole depth.

[0029] S500. Real-time monitoring of the stress and energy evolution of the working face after roof blasting, and then optimizing the design of drilling layout distance.

[0030] In step S500, real-time monitoring includes the following means: (a) Drilled stress gauges are placed on the roadway roof to monitor the expansion of the plastic zone. The crushing zone extends to within 5 m from the roadway wall, and the stress peak shifts to a position ≥ 12 m deep in the ore body. (b) Continuous observation using a microseismic monitoring system, requiring an energy release rate of ≤10 3 J; (c) The stent pressure monitoring system records the cyclic pressure step in real time, and the step shortening amplitude is ≥5%.

[0031] After step S500, the method further includes: S601. Gob-side tunnels shall be supported by a combination of anchor cables and retractable U-shaped steel, with the anchor cable preload being ≥ 200 kN. S602. Add φ153mm pressure relief holes in the base plate with an inclination of -45°, a depth of 8m to 10m, and a spacing of 3m to 5m.

[0032] Example 1 A coal mine with a hard rock roof faced severe stress concentration during mining on one side of its working face. This frequently resulted in tunnel deformation, rock spalling, and roof collapse, posing safety risks such as water inrush, rockburst, and gas outburst. To address this issue, the team decided to employ a deep-hole blasting roof-breaking pressure relief method.

[0033] Preparatory work: Detailed geological data on the mine was collected, including the occurrence of the mine's coal seams and rock formation characteristics. Core sampling experiments were also conducted to obtain drill hole histograms to accurately determine the number, thickness, lithology, and physical and mechanical parameters of the rock layers above the working face's coal seams. Furthermore, data on previous observations of ground subsidence and roof collapse in the goaf was compiled, and analysis determined that the vertical distance H between the critical layer and the mine's coal seams was 30 meters.

[0034] Numerical simulation model development and microseismic monitoring: Based on overburden parameters from the borehole histogram, specialized software was used to construct a numerical simulation model of the working face and one side of the goaf. Combined with data collected by the on-site microseismic monitoring system, the peak location of the working face's advanced abutment pressure, h1, was determined to be 12 meters, the range of influence of the advanced abutment pressure was h2, which was 40 meters. The peak location of the lateral abutment pressure in the goaf, h3, was determined to be 8 meters, and the range of influence of the lateral abutment pressure in the goaf, h4, was determined to be 15 meters.

[0035] Determine the blasting decompression location: Drill hole arrangement in front of the cut: The working face length L = 100m. In the area in front of the cut, drill a blast hole every 10m, for a total of N1 = L / 10 = 10 holes. Drill angle α1 = arctan (H / h1) = arctan (30 / 12) = 68°. Drill depth l1 = H / sinα1 = 30 / sin68° ≈ 32.67m.

[0036] Drilling arrangements along the gob side of the roadway: Drilling holes are arranged along the gob side of the roadway from the peak position h1 of the working face's advanced support pressure, extending outward from the cut to within h2 = 40 m. During excavation, a drill hole is arranged every 15 m along the gob side of the roadway, and during mining, a drill hole is also arranged every 15 m; on the mining side, two drill holes are arranged every 15 m. The distance from the gob side of the roadway to the gob is l = 10 m. The number of drill holes along the gob side is N2 = h2 / 7.5 = 40 / 7.5 ≈ 5.33, so a total of 6 holes are used. The drilling angle α2 = arctan (H / l) = arctan (30 / 10) = 71.57°, and the drilling depth l2 = H / sinα2 = 30 / sin71.57° ≈ 31.83 m. The number of drill holes 1 on the side of the mining slope is N3 = h2 / 15 = 40 / 15 ≈ 2.67, so 3 holes are selected. The angle of drill hole 1 is α3 = arctan (H / h3) = arctan (30 / 8) = 73.3°, and the depth of drill hole 1 is l3 = H / sinα3 = 30 / sin73.3° ≈ 31.53m. The number of drill holes 2 is N4 = N3 = 3 holes. The angle of drill hole 2 is α4 = arctan (H / h4) = arctan (30 / 15) = 63.43°, and the depth of drill hole 2 is l4 = H / sinα4 = 30 / sin63.43° ≈ 33.54m.

[0037] Drilling arrangement for solid coal roadways: Solid coal roadways are arranged outward from the peak position h1 of the working face's advanced support pressure, extending from the cut to within h2 = 40m. Blasting pressure relief holes are arranged every 15m along the working face's inclination along the mining side, with the same layout parameters as those for side drilling holes 2 along the mining side of the gob-side roadway. Furthermore, based on on-site anti-bumping experience, drill holes are arranged every 7.5m in the middle of the roadway along the working face's strike. The number of drill holes, N5 = h2 / 7.5 = 40 / 7.5 ≈ 5.33, is 6; the drilling angle, α5, is 80°, and the drilling depth, l5, is H / sinα5 = 30 / sin80° ≈ 30.77m.

[0038] The arrangement of drilling holes for the mining side of the solid coal roadway meets the following requirements: when the working face advancement speed v>5m / d, the drilling spacing is adjusted to 10m, and α5=80°.

[0039] Drilling and Blasting: Using a suitable drilling rig, drill multiple 80mm diameter holes into the primary control layer according to the locations determined above. Emulsion explosives are charged in an axial continuous coupled manner. The charge is determined based on the hole depth and diameter, and the length of the grouting and sealing section must be no less than one-third of the hole depth. After charging and sealing, blasting is performed according to the blasting design parameters to sever the critical layer.

[0040] Real-time monitoring and optimization: Borehole stress gauge monitoring: Borehole stress gauges are placed on the roadway roof to monitor the expansion range of the plastic zone in real time, ensuring that the crushing zone is controlled within 5m from the roadway wall and the stress peak is successfully transferred to a position ≥12m deep in the coal body.

[0041] Microseismic monitoring system observation: Use microseismic monitoring system for continuous observation, real-time monitoring of electromagnetic radiation intensity, to ensure that its value is stable at ≤10 3 Within the range of J, it reflects that the stress change of the rock formation is in a controllable state.

[0042] Support pressure monitoring: The support pressure monitoring system records the periodic pressure step in real time, and it is observed that the step shortening amplitude is ≥5%, indicating that the blasting pressure relief has a positive effect on the working face pressure.

[0043] Based on these monitoring data, the drilling layout distance is optimized and adjusted to further improve the pressure relief effect.

[0044] Support and bottom plate pressure relief: Gob-side tunnel support: Gob-side tunnels are supported by anchor nets and retractable U-shaped steel. The anchor preload is strictly controlled at ≥200kN to ensure that the tunnel has sufficient support strength, effectively resists the surrounding rock pressure, and ensures tunnel stability.

[0045] Floor relief hole arrangement: φ153mm relief holes were added to the floor, with a -45° inclination, a depth of 9m, and a spacing of 4m. These relief holes effectively relieve stress in the floor rock, preventing deformation problems such as floor heave, and further improving the stress state of the tunnel surrounding rock.

[0046] After implementing this method, problems such as spalling and roof collapse in the working face tunnel were significantly reduced, the stability of the roof was significantly improved, safety hazards such as gas outbursts were effectively controlled, and the working face mining process became safer and more efficient.

[0047] In the description of this specification, the reference terms "one embodiment," "some embodiments," "example," "specific example," or "some examples" mean that the specific features, structures, materials, or characteristics described in conjunction with the embodiment or example are included in at least one embodiment or example of the present invention. In this specification, the schematic representations of the above terms do not necessarily refer to the same embodiment or example. Moreover, the specific features, structures, materials, or characteristics described can be combined in any one or more embodiments or examples in a suitable manner. In addition, those skilled in the art can combine and combine different embodiments or examples described in this specification and features of different embodiments or examples without contradiction.

[0048] The above description is only a preferred embodiment of the present invention and is not intended to limit the present invention. Any modifications, equivalent substitutions, improvements, etc. made within the spirit and principles of the present invention should be included in the scope of protection of the present invention.

Claims

1. A deep hole blasting roof breaking and pressure relief method for lateral stress control in goaf of hard rock mines, characterized in that: The following steps are involved: S100. Based on the mine geological data or the drill hole histogram obtained from the rock coring experiment, the number of layers, thickness, lithology and physical and mechanical parameters of the rock layer above the working face ore body are obtained. Based on the previous observation and analysis of ground settlement and roof collapse in the goaf area, the vertical distance H between the key layer and the ore body is obtained; S200. Based on the overburden parameters obtained from the drilling histogram, a numerical simulation model of the working face and the goaf on one side is established, and the microseismic monitoring data is combined to determine Peak position of leading support pressure of working face h1; Leading support pressure influence range h2; Peak position of lateral support pressure in goaf h3; The influence range of lateral support pressure in goaf h4; S300. Determine the location of blasting decompression in the key layer; S400. According to the position in step S300, multiple boreholes are drilled into the main control layer, and explosives are filled in the boreholes. After sealing the holes, blasting is used to cut off the key layer. S500. Real-time monitoring of the stress and energy evolution of the working face after roof blasting, and then optimizing the design of drilling layout distance.

2. The deep hole blasting and roof breaking pressure relief method for lateral stress regulation in the goaf of hard rock mines according to claim 1 is characterized in that: Step S300 specifically includes the following steps: S301. In front of the cut, the working face is of length L, and blasting holes are drilled every 10m, where: Number of drilling holes: N1=L / 10; Drilling angle: α1=arctanα6; α6=H / h1; Drilling depth: l1=H / sinα1; S302. Along both sides of the empty roadway in the working face advance support pressure peak position h1 outward arrangement, the arrangement range is open cut to the working face advance support pressure influence range h2; During the excavation period, the tunnel is drilled every 15m along the empty side. During the mining period, one drill hole is drilled every 15m. During the mining period, two drill holes are drilled every 15m on the side of the mining area. The distance between the tunnel along the empty side and the goaf is l, where Airside: Number of drilling holes: N2= h2 / 7.5; Drilling angle: α2=arctanα7; α7=H / l; Drilling depth: l2=H / sinα2; Mining side: Number of drilling holes 1: N3 = h2 / 15 Drilling angle 1: α3=arctanα8; α8=H / h3; Drilling depth 1: l3=H / sinα3; Number of drilling holes 2: N4 = N3; Drilling 2 angle: α4=arctanα9; α9=H / h4; Drilling depth 2: l4=H / sinα4; S303. Solid pillar tunnels are arranged outward from the peak position h1 of the working face's advanced support pressure. The layout range is from the opening of the cut to the working face's advanced support pressure influence range h2. Blasting pressure relief holes are drilled every 15 m along the working face's inclination along the mining sidewalls, using the same layout parameters as those for side drilling holes 2 along the gob-side tunnels. In addition, based on field anti-bumping experience, drill holes are arranged every 7.5 m in the middle of the tunnel along the working face's inclination. Number of drilling holes: N5 = h2 / 7.5; Drilling angle: α5=80°; Drilling depth: l5=H / sinα5.

3. The deep hole blasting roof breaking and pressure relief method for lateral stress regulation in hard rock mine goaf according to claim 1 is characterized by: In step S400, the borehole diameter is 76 mm to 89 mm, the hole is charged in an axial continuous coupled manner, the explosive is an emulsion explosive or a water gel explosive, and the length of the grouting sealing section is ≥ 1 / 3 of the hole depth.

4. The deep hole blasting roof breaking and pressure relief method for lateral stress regulation in hard rock mine goaf according to claim 1 is characterized in that: In step S500, the real-time monitoring target is: (a) Drilled stress gauges are placed on the roadway roof to monitor the expansion of the plastic zone. The crushing zone extends to within 5 m from the roadway wall, and the stress peak shifts to a position ≥ 12 m deep in the ore body. (b) Continuous observation using a microseismic monitoring system, requiring an energy release rate of ≤10 3 J; (c) The stent pressure monitoring system records the cyclic pressure step in real time, and the step shortening amplitude is ≥5%.

5. The deep hole blasting roof breaking and pressure relief method for lateral stress control in hard rock mine goaf according to claim 2 is characterized in that: In step S303: The arrangement of drilling holes for the mining side of the solid pillar roadway must meet the following requirements: The drilling spacing is related to the working face advancement speed v. When v>5m / d, the spacing is adjusted to 10m.

6. The deep hole blasting roof breaking and pressure relief method for lateral stress control in hard rock mine goaf according to claim 2 is characterized in that: After step S500, the method further includes: S601. Gob-side tunnels shall be supported by a combination of anchor cables and retractable U-shaped steel, with anchor cable preload ≥ 200 kN. S602. Add φ153mm pressure relief holes in the base plate with an inclination of -45°, a depth of 8m to 10m, and a spacing of 3m to 5m.

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