Method for determining drilling parameters of roof advanced deep hole pre-splitting blasting in coal mining face

Abstract

The application discloses a method for determining parameters of advanced deep-hole pre-splitting blasting holes of a coal mining face roof, and belongs to the technical field of coal mine safety mining. The method comprises the following steps: obtaining basic technical parameters of the coal mining face; calculating a theoretical rock stratum caving zone thickness; presetting the number of drilling holes as N, and calculating parameters of each blasting hole; determining a maximum hole spacing Lhole between adjacent blasting holes, and calculating a blasting fissure zone radius R according to physical and mechanical parameters of explosives and rocks; comparing L with 2R to determine whether a pressure relief blind area exists, and if the pressure relief blind area exists, increasing the number of drilling holes to N+1, continuously calculating parameters of each blasting hole, and outputting the parameters of the blasting holes after the pressure relief blind area no longer exists. The application changes a traditional drawing method into a parameterized automatic calculation method, and through built-in blind area checking and self-adaptive adjustment closed loop, the design efficiency and the roof treatment reliability are significantly improved.

Description

Technical Field

[0001] This invention relates to the field of coal mine safety mining technology, and in particular to a method for determining drilling parameters for advanced deep-hole pre-splitting blasting of the roof of a coal mining face. Background Technology

[0002] In coal mining, when the roof of the coal seam is composed of hard rock strata (such as sandstone and limestone), the roof is not prone to natural collapse, resulting in a large area of ​​overhang. As the working face advances, the overhang area increases, which may trigger strong initial or periodic pressure, leading to support overload, severe roadway deformation, or even roof collapse accidents. To control roof collapse, advanced deep-hole pre-splitting blasting technology is often used in engineering. This involves drilling holes of a certain depth and angle into the roof in the two roadways or cut-outs of the working face, and then using blasting to pre-create fissures in the roof strata, thereby forcing the roof to collapse according to the design.

[0003] In existing technologies, the determination of blasting hole parameters (such as the number of boreholes, elevation angle, and length) mainly relies on drafting methods. Specifically, technicians manually draw rays at different angles representing boreholes, starting from the drilling site location, using CAD or other drafting software based on the mine's geological profile. The angles and lengths are repeatedly adjusted by measuring the coordinates of the intersection points of these rays with the target roof layer until design constraints (such as maximum borehole length limits and the area of ​​roof treatment coverage) are met. This method is cumbersome, requiring redrawing and remeasurement for each parameter adjustment, resulting in extremely low efficiency. The workload increases exponentially when comparing 3-hole and 4-hole schemes or verifying the existence of blasting blind zones.

[0004] Therefore, there is an urgent need for a method that can quickly, accurately, and adaptively determine the parameters of deep hole pre-fracture blasting holes in the roof to solve the problem of low efficiency in existing technologies. Summary of the Invention

[0005] To address this, the present invention proposes a method for determining the parameters of pre-splitting blasting holes in the roof of a coal mining face. This method overcomes the efficiency deficiencies of traditional drawing methods through parametric calculation and automatic closed-loop verification, enabling rapid and accurate determination of blasting hole parameters.

[0006] To address the aforementioned technical problems, the present invention provides the following technical solution: A method for determining the parameters of advanced deep-hole pre-splitting blasting holes in the roof of a coal mining face includes the following steps: Step S1: Obtain the basic technical parameters of the coal mining face. The basic technical parameters include at least: coal seam thickness M, design recovery rate r, average fracture and swelling coefficient of overlying coal and rock strata k, vertical distance of roof treatment layer H, roadway width d, roadway height h1, coal seam dip angle α, and maximum blasting hole length L. max And the physical and mechanical parameters of the explosive and the rock; Step S2: Calculate the thickness h of the rock strata collapse zone under theoretical conditions using the formula h=(M×r) / (k-1); Step S3: Preset the number of boreholes to N, and calculate the parameters of the N blasting holes in the upper and lower roadway drilling sites respectively. The parameters include at least the horizontal rotation angle, elevation angle and hole length of each blasting hole. Step S4: Based on the blasting hole parameters calculated in step S3, determine the maximum hole spacing L between adjacent blasting holes. 孔距 ; Step S5: Calculate the radius R of the fracture zone formed after the blasting hole is blasted, based on the physical and mechanical parameters of the explosive and the rock. Step S6: Compare the maximum hole spacing L 孔距 With respect to the radius R of the fractured region: If L 孔距 If ≤2R, then the blast hole parameters calculated in step S3 are output as the final design parameters. If L 孔距 If the value is greater than 2R, it is determined that there is a pressure relief blind zone, and the number of boreholes is adjusted to N+1. The parameters of the blasting holes in the upper and lower roadway drilling sites are recalculated, and the process proceeds to step S4.

[0007] In some embodiments of the present invention, the number of boreholes in step S3 is preset to 3. For the three blasting holes S1, S2, and S3 in the upper roadway drilling site, the parameters of the blasting holes are obtained using the following method: The horizontal rotation angle of the blasting hole S1 is 90°, the elevation angle is β1, and the hole length L1 is calculated according to the formula: L1 = h × (1 + (H - (d × tanα + h1) × cosα) / h) / sin(β1 + α) is calculated. The length of the blast hole S3 is L max The horizontal rotation angle is 90°, and its elevation angle β3 is calculated according to the formula: β3=arcsin(h×(1+(H-(d×tanα+h1)×cosα) / h) / L max )-α calculation; The elevation angle β2 of the blasting hole S2 is the midpoint between the elevation angles of S1 and S3, and its hole length L2 is calculated based on the sine relationship corresponding to the midpoint.

[0008] In some embodiments of the present invention, for the N blasting holes S1, S2...S in the upper haulage drilling site N And N>3, the parameters of the blast hole are obtained using the following method: The horizontal rotation angle of the blasting hole S1 is 90°, the elevation angle is β1, and the hole length L1 is calculated according to the formula: L1 = h × (1 + (H - (d × tanα + h1) × cosα) / h) / sin(β1 + α) is calculated. Explosion Hole S N The length of the hole is L max The horizontal rotation angle is 90°, and its elevation angle β N According to the formula: β N =arcsin(h×(1+(H-(d×tanα+h1)×cosα) / h) / L max )-α calculation; The blast holes S2 to S2 are located in the middle. N The elevation angle will be between S1 and S N The difference in elevation angle is divided into N-1 equal parts, and the length of each hole is calculated based on the sine relationship corresponding to its respective elevation angle.

[0009] In some embodiments of the present invention, in step S4, the maximum hole spacing L between adjacent blasting holes 孔距 Located at hole S N-1 With Kong S N Between, and calculated according to the following formula: The elevation angle difference η between adjacent blasting holes is calculated according to the following formula: η=(β1-β N ) / (N-1) Explosion Hole S N-1 elevation angle β N-1 Calculate according to the following formula: β N-1 =β N +η Explosion Hole S N-1 The angle γ between the coal seam and the coal seam plane is calculated using the following formula: γ=β N-1 -α Then the blast hole S N-1 Length L N-1 Calculate according to the following formula: L N-1 =h / sinγ Maximum hole spacing L between adjacent blasting holes 孔距 Calculate according to the following formula: L 孔距 =L N-1 *sinβ N-1 .

[0010] In some embodiments of the present invention, calculating the fracture zone radius R in step S5 includes: According to formula P d =(ρ e ×D 2 / 8)×(dc / d b ) 6 ×n calculates the initial impact pressure acting on the borehole wall, where ρ e Let d be the density of the explosive, D be the detonation velocity, and d be the detonation velocity. c d is the diameter of the explosive. b Where is the borehole diameter, and n is the stress amplification factor; According to the formula R=((P) d / σ t )^(1 / a))×r b Calculate the radius R of the fractured region, where σ t Let be the tensile strength of the rock, 'a' be the attenuation exponent, and 'r' b The radius of the borehole is given.

[0011] In some embodiments of the present invention, the stress amplification factor n ranges from 8 to 12.

[0012] In some embodiments of the present invention, the attenuation index a ranges from 1.25 to 1.35.

[0013] In some embodiments of the present invention, step S8 is further included: according to the hole length L of each blasting hole N According to the sealing length L 封 =ceil(L N / 3) Calculate the sealing length of each hole, where ceil is the round-up function.

[0014] In some embodiments of the present invention, based on the charge length L 药 =L N -L 封 Calculate the charge amount for each hole. The charge amount is calculated based on the product of the charge length and the charge amount per unit length. The charge amount per unit length is determined based on the type of explosive used.

[0015] In some embodiments of the present invention, the sealing hole length L 封 The calculation can also be determined by rounding to the nearest whole number or rounding down and then adding a preset safety margin, which is 1 to 2 meters.

[0016] The technical solution of the present invention has the following technical effects compared with the prior art: The method for determining parameters of pre-fracturing blasting holes in the roof of a coal mining face provided by this invention transforms the traditional drawing method into a parametric automatic calculation method. Through built-in blind zone verification and adaptive adjustment closed-loop, it significantly improves design efficiency and roof treatment reliability. The entire process requires no manual intervention, solving the technical problems of low efficiency, reliance on experience, and difficulty in adaptive optimization inherent in traditional drawing methods. Attached Figure Description

[0017] The preferred embodiments of the present invention will be described in detail below with reference to the accompanying drawings, which will help to understand the purpose and advantages of the present invention, wherein: Figure 1 This is a flowchart of the method for determining the parameters of the pre-fracture blasting hole in the roof of a coal mining face according to the present invention. Figure 2 This is a schematic diagram of the arrangement of three sets of deep pre-splitting blasting holes in the roof of a coal mining face. Figure 3 This is a schematic diagram of the arrangement of three sets of deep pre-splitting blasting holes in the roof of a coal mining face. Figure 4 for Figure 2 This is a schematic diagram showing the arrangement of three sets of pre-splitting blasting holes in the roof of a coal mining face. Detailed Implementation

[0018] The technical solution of the present invention will now be clearly and completely described with reference to the accompanying drawings. Obviously, the described embodiments are only some, not all, of the embodiments of the present invention. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0019] In the description of this invention, it should be noted that the terms "center," "upper," "lower," "left," "right," "vertical," "horizontal," "inner," and "outer," etc., indicate the orientation or positional relationship based on the orientation or positional relationship shown in the accompanying drawings. They are used only for the convenience of describing the invention and for simplifying the description, and do not indicate or imply that the device or element referred to must have a specific orientation, or be constructed and operated in a specific orientation. Therefore, they should not be construed as limitations on the invention. Furthermore, the terms "first," "second," and "third" are used for descriptive purposes only and should not be construed as indicating or implying relative importance.

[0020] In the description of this invention, it should be noted that, unless otherwise explicitly specified and limited, the terms "installation," "connection," and "joining" should be interpreted broadly. For example, they can refer to a fixed connection, a detachable connection, or an integral connection; they can refer to a direct connection or an indirect connection through an intermediate medium; and they can refer to the internal communication between two components. Those skilled in the art can understand the specific meaning of the above terms in this invention based on the specific circumstances.

[0021] Furthermore, the technical features involved in the different embodiments of the present invention described below can be combined with each other as long as they do not conflict with each other.

[0022] like Figure 1The illustration shows a specific embodiment of the method for determining the parameters of advanced deep-hole pre-splitting blasting holes in the roof of a coal mining face according to the present invention. This method includes steps for obtaining basic parameters, calculating the theoretical strata collapse zone thickness, calculating the parameters of a preset number of boreholes, determining the maximum hole spacing, calculating the radius of the fracture zone, and identifying and adaptively adjusting blind zones. These steps achieve a fully automated closed-loop design, eliminating the subjectivity of judging blind zones based on experience, and technically ensuring the integrity and safety of roof treatment.

[0023] Specifically, at the coal mine site, technicians first obtain the following basic technical parameters through the mine geological report, borehole columnar section, roadway measured data, and explosive performance specifications: Coal seam thickness M: the true thickness of the pure coal seam excluding interbedded rock; design recovery rate r (dimensionless, usually taken as 0.80–0.95); average fracture coefficient k of the overlying coal and rock strata (dimensionless, empirical value 1.25–1.35); vertical distance H of the roof treatment layer: the distance vertically upward from the roadway floor (or drilling site reference point) to the top surface of the target roof treatment layer; roadway width d; roadway height h1; coal seam dip angle α; maximum blasting hole length L. max The depth of penetration is determined by the maximum drilling depth and the explosive loading capability; as well as the physical and mechanical parameters of the explosive and the rock, including the explosive density ρ. e Detonation velocity D, explosive diameter d _c , borehole diameter d _b Rock tensile strength σ _t The decay index a, Poisson's ratio μ, etc.

[0024] Step S2: Obtain the theoretical thickness h of the caving zone based on the following formula: h=(M×r) / (k-1) Among them, the fragmentation coefficient k reflects the characteristic of the volume increase after the rock is broken, and the theoretical rock stratum collapse zone thickness h ensures that the collapse zone can effectively fill the goaf after blasting and avoid roof suspension.

[0025] Step S3: The number of boreholes is preset to N, and in this implementation scheme, N is preferably 3. The first set of preset formulas calculates the horizontal rotation angle (+90° for the upper roadway and -90° for the lower roadway), elevation angle, and borehole length for each borehole in both the upper and lower roadway drilling areas based on the geometric relationship of the fan-shaped arrangement. For example... Figure 2 As shown, the first set of preset formulas takes the drilling site location as the origin and the roof treatment layer as an oblique line (considering the coal seam dip angle) and establishes an accurate mapping between the borehole length and the elevation angle through trigonometric functions.

[0026] Specifically, the boreholes in the upper roadway drilling site are designated as S1, S2, and S3, with a fixed horizontal rotation angle of 90° (perpendicular to the roadway axis pointing towards the goaf). The maximum allowable elevation angle is set to β (determined by the drilling rig performance, typically 45°–60°).

[0027] The elevation angle of hole S1 is taken as the maximum value β1, and the hole length L1 is calculated by the following formula: L1=h×(1+(H-(d×tanα+h1)×cosα) / h) / sin(β1+α) exist Figure 2 In the equation, the hole length L1 is the length of the diagonal line from the drilling site location (the top corner of the tunnel) to the upper boundary of the roof treatment layer. The (H-(d×tanα+h1)×cosα) in the numerator corrects for the influence of the tunnel cross-sectional dimensions on the vertical offset of the target layer.

[0028] The hole length of S3 is the maximum allowable value L of the equipment. max The elevation angle β3 is calculated by the following formula: β3=arcsin(h×(1+(H-(d×tanα+h1)×cosα) / h) / L max )–α This formula ensures that, under the maximum hole length limit, the drill hole can reach just the shallowest part of the top plate treatment layer.

[0029] The elevation angle of hole S2 is taken as the arithmetic mean of the elevation angles of S1 and S3: β2 = (β1 + β3) / 2; The hole length L2 is calculated by substituting β2 into the same functional form as L1: L2=h×(1+(H-(d×tanα+h1)×cosα) / h) / sin(β2+α) For the lower haulage drilling site, such as Figure 2 As shown, the boreholes are numbered N1, N2, and N3. Symmetrical treatment: the horizontal rotation angle is -90°, and "+α" in all formulas is replaced with "-α".

[0030] By using arithmetic equal division of the elevation angle, the projected spacing of the boreholes on the top plate treatment layer is made as uniform as possible, thereby minimizing the maximum hole spacing L. 孔距 This facilitates subsequent verification. The formula incorporates the effects of roadway dimensions (d, h1) and coal seam dip angle (α), adapting to different mine conditions and avoiding the hassle of manual correction based on roadway cross-sections required in traditional methods.

[0031] Step S4: Based on the calculated N drilling parameters, extract the projection points of adjacent boreholes on the top plate treatment layer, and calculate the spatial straight-line distance between the two points, which is the borehole spacing. This L... 孔距 This characterizes the width of the area that may exist between the fan-shaped holes but is not covered by blasting. The maximum hole spacing L is... 孔距 Then located at S N-1 With S N Between (e.g.) Figure 2 , Figure 3(as shown), and the maximum hole spacing is calculated using the following formula: The difference in elevation angle η between adjacent blast holes is calculated according to the following formula: η=(β1-β N ) / (N-1) Explosion Hole S N-1 elevation angle β N-1 Calculate according to the following formula: β N-1 =β N +η Explosion Hole S N-1 The angle γ between the plane of the coal seam and the coal seam plane is calculated according to the following formula: γ=β N-1 -α Then the blast hole S N-1 Length L N-1 Calculate according to the following formula: L N-1 =h / sinγ Maximum hole spacing L between adjacent blasting holes 孔距 Calculate according to the following formula: L 孔距 =L N-1 *sinβ N-1 .

[0032] Specifically, in a 3-hole fan-shaped arrangement, the maximum hole spacing usually occurs between the middle hole (S2) and the hole with the maximum elevation angle (S3). Here, the elevation angle of blasting hole S2 is β2 = (β1 + β3) / 2; the angle between blasting hole S2 and the coal seam plane is γ = β2 - α; the length of blasting hole S2 is L2 = h / sinγ; and the hole spacing between blasting holes S2 and S3, i.e., the maximum hole spacing L... 孔距 =L2*sinβ N-1 .

[0033] Step S5: Based on the physics of explosive detonation and the mechanics of rock fracture, first calculate the initial impact pressure P on the borehole wall under the condition of uncoupled charge. d Then, combining the tensile strength of the rock and the stress wave attenuation law, the radius R of the fracture zone after a single-hole blast is calculated. Within this radius R, the rock mass is completely broken into blocks and loses its load-bearing capacity.

[0034] Specifically, for decoupled charge blasting, after the explosive detonates, the blast gas expands and fills the borehole, and the initial radial stress peak P acting on the borehole wall... d Calculate using the following formula: P d =(ρ e ×D 2 / 8)×(d c / d b ) 6 ×n In the formula, n is the stress increase factor when the explosive gas collides with the rock wall, which is taken as 8 to 12. (d) c / d b ) 6 It reflects the dramatic effect of the decoupling coefficient on the pressure, and the index 6 comes from the superposition of multiple reflections of the detonation wave in the gap.

[0035] The radius R of the fractured region is: R=((P d / σ t )^(1 / a))×r b Where 'a' is the stress wave attenuation index, which is typically 1.25-1.35 for columnar drug packs; in this embodiment, 1.3 is selected; r b =d b / 2 represents the borehole radius.

[0036] When the impact pressure decays to equal the tensile strength σ of the rock t When the rock undergoes tensile fracture, this distance is the boundary of the fracture zone. In field applications, simply changing the explosive charge to a different diameter or adjusting the borehole diameter will automatically update the R value, thus affecting the blind zone assessment results. This provides a theoretical basis for optimizing blasting techniques.

[0037] Step S6: Place L 孔距 Compare with 2R. If L 孔距 If ≤2R, then the fracture zones of adjacent holes overlap or are exactly connected, there is no pressure relief blind zone, the current number of boreholes meets the requirements, and the parameters are output directly.

[0038] If L 孔距 If the value is greater than 2R, it indicates that there is an intact rock bridge that has not been destroyed between adjacent fracture zones, i.e., a "pressure relief blind zone". In this case, the number of boreholes must be increased (usually by 1, i.e., becoming N+1) to reduce the spacing between boreholes and eliminate the blind zone. When the number of boreholes is adjusted to N+1, the system automatically calls the second set of preset formulas, recalculates the parameters of all boreholes, and outputs the final design results.

[0039] For example, when the number of holes is adjusted to 4, such as Figure 3 As shown, the boreholes are numbered S1, S2, S3, and S4, and the horizontal rotation angle remains 90°.

[0040] S1 hole: It is exactly the same as S1 in the 3-hole scheme, with elevation angle = β1 and hole length L1 formula unchanged.

[0041] S4 hole: exactly the same as S3 in the 3-hole scheme, hole length = L max The formula for calculating the elevation angle β4 is the same.

[0042] S2 and S3 holes: Divide the difference in elevation angle (β1-β4) between S1 and S4 into three equal parts, then: β2=β4+(β1-β4) / 3, β3=β4+2×(β1-β4) / 3.

[0043] Substituting the hole lengths L2 and L3 into their respective elevation angles, the following formula applies: L N =h×(1+(H-(d×tanα+h1)×cosα) / h) / sin(β N Calculate using +α).

[0044] For the lower haulage drilling site, such as Figure 3 As shown, the boreholes are numbered N1, N2, N3, and N4, and are treated symmetrically: the horizontal rotation angle is -90°, and "+α" in all formulas is replaced with "-α". The 4-hole scheme is a remedial measure that automatically switches after a blind zone is found during the 3-hole verification. By dividing the elevation angle interval into N-1 equal parts, the projection points of the N boreholes on the top plate treatment layer are ensured to be equidistantly distributed. The maximum hole spacing L is significantly reduced compared to the 3-hole scheme, which can effectively eliminate the blind zone, and the increase in the number of boreholes is limited, making it economical and reasonable.

[0045] The above steps enabled a closed-loop design for determining blasting hole parameters, avoiding manual iterations and reducing design time from hours to minutes. Simultaneously, by directly linking the blasting mechanism with drilling layout parameters, the subjectivity of relying on experience-based judgments was eliminated, technically ensuring the integrity and safety of the roof treatment.

[0046] For any given blast hole, its total length L N It consists of a sealing section (without explosives near the borehole opening, used to fill the hole with stemming material) and a charging section. To ensure that the blasting energy effectively acts deep into the roof while preventing borehole perforation, the sealing length is L. 封 Calculate using the following formula: L 封 =ceil(L N / 3) Where ceil represents the floor function (e.g., 17.43 rounded up to 18). Then, the charge length L... 药 =L N -L 封 .

[0047] Using one-third of the borehole length as the sealing length is an empirically optimal ratio verified by extensive coal mine practice. This ensures sufficient filling length to seal off explosive gases without wasting drilling resources. Rounding up results in a sealing length slightly longer than the theoretical value, increasing safety redundancy. This method enables fully automated calculation of sealing parameters, eliminating the need for manual estimation.

[0048] Besides the rounding up method, another method can be used: Rounding to the nearest integer: L 封 =round(L N / 3), such as 17.43→17m.

[0049] Rounding down and adding a safety margin: L 封 =floor(L N / 3)+Δ, where Δ is the preset safety margin, which is 1 to 2m. For example, for gas outburst mines, a larger safety margin (Δ=2) is preferred.

[0050] For those skilled in the art, the sealing length can be flexibly selected based on two methods, according to factors such as mine safety management requirements, the stability of the mine's roof, and gas levels.

[0051] Based on the above charge length L 药 The charge amount for each hole can be calculated. The charge amount is calculated based on the product of the charge length and the charge amount per unit length, which is determined according to the type of explosive used.

[0052] Specifically, the charge per unit length q (kg / m) is pre-determined based on the type of explosive selected. For commonly used permissible explosives in coal mines: 94mm large diameter cartridge: q≈5.7kg / m 75mm standard cartridge: q≈2.8kg / m Then the charge per hole Q = q × L 药 .

[0053] The system directly links the charge quantity calculation to the type of explosive, making it easier for on-site construction personnel to requisition explosives according to the design. Simultaneously, the system can automatically calculate blasting costs based on the charge quantity, aiding in economic comparisons.

[0054] Obviously, the above embodiments are merely illustrative examples for clear explanation and are not intended to limit the implementation. Those skilled in the art will recognize that other variations or modifications can be made based on the above description. It is neither necessary nor possible to exhaustively list all possible implementations here. However, obvious variations or modifications derived therefrom are still within the scope of protection of this invention.

Claims

1. A method for determining the parameters of advanced deep-hole pre-splitting blasting holes in the roof of a coal mining face, characterized in that, Includes the following steps: Step S1: Obtain the basic technical parameters of the coal mining face. The basic technical parameters include at least: coal seam thickness M, design recovery rate r, average fracture and swelling coefficient of overlying coal and rock strata k, vertical distance of roof treatment layer H, roadway width d, roadway height h1, coal seam dip angle α, and maximum blasting hole length L. max And the physical and mechanical parameters of the explosive and the rock; Step S2: Calculate the thickness h of the rock strata collapse zone under theoretical conditions using the formula h=(M×r) / (k-1); Step S3: Preset the number of boreholes to N, and calculate the parameters of the N blasting holes in the upper and lower roadway drilling sites respectively. The parameters include at least the horizontal rotation angle, elevation angle and hole length of each blasting hole. Step S4: Based on the blasting hole parameters calculated in step S3, determine the maximum hole spacing L between adjacent blasting holes. 孔距 ; Step S5: Calculate the radius R of the fracture zone formed after the blasting hole is blasted, based on the physical and mechanical parameters of the explosive and the rock. Step S6: Compare the maximum hole spacing L 孔距 With respect to the radius R of the fractured region: If L 孔距 If ≤2R, then the blast hole parameters calculated in step S3 are output as the final design parameters. If L 孔距 If the value is greater than 2R, it is determined that there is a pressure relief blind zone, and the number of boreholes is adjusted to N+1. The parameters of the blasting holes in the upper and lower roadway drilling sites are recalculated, and the process proceeds to step S4.

2. The method according to claim 1, characterized in that, In step S3, the number of boreholes is preset to 3. For the three blasting holes S1, S2, and S3 in the upper roadway drilling site, the parameters of the blasting holes are obtained using the following method: The horizontal rotation angle of the blasting hole S1 is 90°, the elevation angle is β1, and the hole length L1 is calculated according to the formula: L1 = h × (1 + (H - (d × tanα + h1) × cosα) / h) / sin(β1 + α) is calculated. The length of the blast hole S3 is L max The horizontal rotation angle is 90°, and its elevation angle β3 is calculated according to the formula: β3=arcsin(h×(1+(H-(d×tanα+h1)×cosα) / h) / L max )-α calculation; The elevation angle β2 of the blasting hole S2 is the midpoint between the elevation angles of S1 and S3, and its hole length L2 is calculated based on the sine relationship corresponding to the midpoint.

3. The method according to claim 1, characterized in that, For the N blasting holes S1, S2...S in the upper haulage drilling site N And N>3, the parameters of the blast hole are obtained using the following method: The horizontal rotation angle of the blasting hole S1 is 90°, the elevation angle is β1, and the hole length L1 is calculated according to the formula: L1 = h × (1 + (H - (d × tanα + h1) × cosα) / h) / sin(β1 + α) is calculated. Explosion Hole S N The length of the hole is L max The horizontal rotation angle is 90°, and its elevation angle β N According to the formula: β N =arcsin(h×(1+(H-(d×tanα+h1)×cosα) / h) / L max )-α calculation; The blast holes S2 to S2 are located in the middle. N The elevation angle will be between S1 and S N The difference in elevation angle is divided into N-1 equal parts, and the length of each hole is calculated based on the sine relationship corresponding to its respective elevation angle.

4. The method according to claim 1, characterized in that, In step S4, the maximum hole spacing L between adjacent blasting holes 孔距 Located at hole S N-1 With Kong S N Between, and calculated according to the following formula: The elevation angle difference η between adjacent blasting holes is calculated according to the following formula: η=(β1-β N ) / (N-1) Explosion Hole S N-1 elevation angle β N-1 Calculate according to the following formula: b N-1 =b N +n Explosion Hole S N-1 The angle γ between the coal seam and the coal seam plane is calculated using the following formula: c=b N-1 -a Then the blast hole S N-1 Length L N-1 Calculate according to the following formula: L N-1 =h / sinγ Maximum hole spacing L between adjacent blasting holes 孔距 Calculate according to the following formula: L 孔距 =L N-1 *sinβ N-1 。 5. The method according to claim 1, characterized in that, The calculation of the fracture zone radius R in step S5 includes: According to formula P d =(ρ e ×D 2 / 8)×(d c / d b ) 6 ×n calculates the initial impact pressure acting on the borehole wall, where ρ e Let d be the density of the explosive, D be the detonation velocity, and d be the detonation velocity. c d is the diameter of the explosive. b Where is the borehole diameter, and n is the stress amplification factor; According to the formula R=((P) d / σ t )^(1 / a))×r b Calculate the radius R of the fractured region, where σ t Let be the tensile strength of the rock, 'a' be the attenuation exponent, and 'r' b The radius of the borehole is given.

6. The method according to claim 5, characterized in that, The stress amplification factor n ranges from 8 to 12.

7. The method according to claim 5, characterized in that, The attenuation index α ranges from 1.25 to 1.

35.

8. The method according to any one of claims 1-7, characterized in that, It also includes step S8: based on the hole length L of each blasting hole N According to the sealing length L 封 =ceil(L N / 3) Calculate the sealing length of each hole, where ceil is the round-up function.

9. The method according to claim 8, characterized in that, Based on the charge length L 药 =L N -L 封 Calculate the charge amount for each hole. The charge amount is calculated based on the product of the charge length and the charge amount per unit length. The charge amount per unit length is determined based on the type of explosive used.

10. The method according to any one of claims 1-7, characterized in that, The sealing length L 封 The calculation can also be determined by rounding to the nearest whole number or rounding down and then adding a preset safety margin, which is 1 to 2 meters.

Images