Large-dip-angle fully mechanized caving face crossheading asymmetric supporting method

Abstract

The invention belongs to the technical field of coal mining equipment, and particularly relates to a large-dip-angle fully mechanized caving face crossheading asymmetric supporting method. Comprising the following steps: S1, calculating a stable safe span of a rock stratum; s2, calculating a roadway span; s3, determining coal mine crossheading roof stability classification indexes; s4, determining the basic form and main parameters of crossheading support; s5, according to the stability of the top plate, the crossheading is subjected to regional division, supporting is conducted in a segmented mode according to the corresponding supporting mode, and when the top plate with poor stability is encountered, an asymmetric supporting mode is selected for supporting; s6, adopting an anchor rod and anchor cable combined supporting mode for asymmetric supporting of the gate road for medium-stable or unstable top plates, and adopting a full anchor cable supporting mode for asymmetric supporting of the gate road for extremely-unstable top plates; s7, calculating related parameters of the anchor rod support; s8, calculating anchor cable support related parameters; and S9, selecting a proper number of anchor rods and anchor cables to carry out corresponding asymmetric support on the crossheading.

Description

Technical Field

[0001] The invention belongs to the technical field of coal mining equipment, and in particular relates to an asymmetric supporting method for a large-angle fully-mechanized caving working face drift. Background Art

[0002] Coal mine tunnel support has evolved from the initial wooden support to masonry support and steel support, and has developed to the most widely used anchor support. Anchor support is a safe, effective and economical support technology for tunnels, and is one of the key technologies to achieve high-yield and efficient production of the working face.

[0003] Different from the normal fully mechanized caving working face, the high-angle working face refers to the working face when the inclination of the ore layer is greater than 45 degrees. The stability of the mine wall is poor, and it is prone to problems such as collapse and landslide, and the mining risk is high. Therefore, the surrounding rock has higher and higher requirements on the stability of the support system, so anchor support is adopted.

[0004] The following factors should be considered when carrying out anchor support: (1) The anchor support system can provide higher axial and lateral restraint forces, and the top plate of the anchoring area has higher strength and rigidity, which effectively restrains the destruction and deformation of the surrounding rock of the top plate and limits the sinking of the upper top plate. (2) The top plate anchors are formed into a whole through the metal mesh to prevent the broken top plate from falling and jointly restrain the deformation and destruction of the top plate in the anchoring area. (3) When using prestressed anchor cables for reinforced support, the anchor cables suspend the top plate of the lower anchoring area in the upper stable rock layer and support together with the anchor rods, which can increase the stability of the top plate and ensure the safety and reliability of the tunnel top plate. (4) The top plate adopts high-strength anchor support, which can increase the spacing between anchor rods, reduce the density of anchor support, and increase the tunnel excavation speed. (5) The top plate adopts high-strength anchor support to reduce the pressure of the coal body on both sides and improve the overall stability of the tunnel top plate and coal side.

[0005] In the process of support, the overall form of fully symmetrical support is often adopted. However, due to different conditions, asymmetric phenomena may occur in some sections and they are extremely unstable. According to the different support strengths, the distribution morphology of the plastic zone of the surrounding rock of the tunnel in different sections, the vertical stress distribution characteristics, the horizontal stress distribution characteristics and the influence of the deformation of the tunnel surrounding rock should be analyzed. Therefore, in the asymmetric area of ​​the tunnel, the combined support method of anchor rods and anchor cables is adopted when the roof is relatively stable, and the full anchor cable support method is used when it is relatively unstable. Summary of the invention

[0006] In order to solve the problems of difficult support, unstable support and coal seam collapse in the mining process caused by an asymmetric drift, the present invention provides an asymmetric support method for a drift of a large-angle fully-mechanized caving working face.

[0007] The present invention adopts the following technical scheme: an asymmetric support method for a large-angle fully-mechanized caving working face heading, comprising: S1: Calculate the safe span of rock stability, including mining influence coefficient, burial depth influence coefficient, ground stress influence coefficient, rock mass strength influence, and rock integrity influence; S2: Calculate the tunnel span; S3: According to the distance S between the hard rock layer and the roof, the stable span L of the hard rock layer and the change of the tunnel span B, the classification index of the stability of the coal mine drift roof is determined; S4: According to the actual data measurement of the coal mine, the basic form and main parameters of the drift support are determined; S5: Divide the drift into areas according to the stability of the roof, and support them in sections according to the corresponding support methods. When encountering a roof with poor stability, choose an asymmetric support method for support; S6: For moderately stable or unstable roofs, the asymmetric support of the trench adopts the combined support of anchor rods and cables; for extremely unstable roofs, the asymmetric support of the trench adopts the full anchor cable support; S7: Calculate the parameters related to anchor support; S8: Calculate the parameters related to anchor cable support; S9: Select the appropriate number of anchor rods and cables to provide corresponding asymmetric support for the drift.

[0008] Step S1 includes: When the thickness of the immediate roof is 4 times the mining height, the mining concentration coefficient is the smallest. When the tunnel is affected by mining once, the mining concentration coefficient is taken as 2.5, and the influence coefficient when affected by mining once is determined as: Determine the influence coefficient of burial depth as: For the same mine, the distribution of ground stress is basically the same, and the influence coefficient of ground stress on tunnel stability is: The rock mass strength is very different from the rock strength measured in the laboratory. Therefore, in the process of roof stability classification, it is necessary to consider the strength difference between rock and rock mass and determine that the rock mass strength is 0.65 times the rock strength, that is, The influence coefficient of rock formation integrity is determined as: Where: —Thickness of rock formations; —Effective thickness of rock formation; Finally, the safe span for the rock formation to maintain its stability is obtained: .

[0009] In step S2, the stable span B of the roof rock layer is: Where: —tunnel span, m; —Stable distance from the top plate, m.

[0010] —The collapse angle of the tunnel.

[0011] In step S3, considering the main influencing factors of roof stability, according to the distance s between the hard rock layer and the roof, the stable span L of the hard rock layer and the change of the tunnel span B, the roof stability classification index of the drift is determined as follows: Class II: Stable, m has stable rock formations, ; Class III: moderately stable, m has stable rock formations, ; Category IV: Unstable, m has stable rock formations, ; Class V: Extremely unstable, m has stable rock formations, .

[0012] In step S7, the required anchor support parameters are calculated using the following formula: (1) Determination of anchor rod length L=L1+L2+L3 Where: L—anchor length, m; L1—anchor exposed length, m; L2—anchor effective length, m; L3—anchor anchor length, m. Determination of anchor exposed length L1: L1 = pad thickness + nut thickness + (0.02 ~ 0.03) m Determination of the effective length L2 of the anchor rod: There are three methods to determine the effective length of the anchor rod. One is to use the sonic method to measure the loosening range of the surrounding rock of the tunnel; the second is to use a rock formation detection analyzer for measurement; the third is to use the Pugh natural equilibrium arch theory to determine it; The effective length L2 of the tunnel top anchor is calculated as follows: The effective length L2 of the tunnel side anchor is calculated as follows: Where: f—Praeter coefficient; B—tunnel span, m; H—tunnel excavation height, m; φ—internal friction angle, °; (2) Determination of spacing between anchor rods When supporting the roadway with anchor rods, the construction technology is usually taken into consideration and the spacing between rows is equal. The spacing between anchor rods D is calculated as follows: D≤0.5L~0.6L For tunnel support, considering the influence of advance support pressure, the spacing between anchor bolts can be appropriately reduced; (3) Determination of anchor rod diameter The anchor diameter d is usually calculated as follows: In order to increase the tunnel support strength, the anchor diameter can be appropriately increased; (4) Determination of anchoring force of anchor bolt Anchoring force of top anchor: Q = 3 × 1.32 × 1 × 1.5 = 5.94t Anchoring force of anchor rod: Q = 3 × 1.05 × 1 × 1.2 = 3.78 t The anchoring force of the top anchor rod is Q≥5.94t, and the anchoring force of the auxiliary anchor rod is Q≥3.78t.

[0013] In step S8, the anchor cable support parameters are calculated using the formula, and the specific calculation formula is as follows: (1) Anchor cable length According to the design principle of anchor cable support, the length of the anchor cable depends on the lithology of the roadway roof, the rock mass structure, the size of the roadway project and the position of the stable rock layer. The anchor cable anchoring section should be designed within 1.0m in the stable rock layer. According to the lithology analysis of the working face, the anchor cable L 锚索 =L 外露 +L 顶煤 +L 稳定岩层 Where: L anchor cable - anchor cable length, unit: m; L exposed - exposed length of anchor cable; L top coal - arranged along the roof of coal seam; L hard rock - length of stable rock layer of anchor cable anchoring section, take 1.0m; (2) Spacing between anchor cables The spacing between anchor cables is determined based on the dangerous rock load with potential collapse tendency that the anchor cables need to bear when the anchor rod fails.

[0014] ① Natural equilibrium height b of the tunnel; b=a / fk Where: a is half of the tunnel excavation width, in m; fk is the Proctor coefficient of the roof rock; ② Potential falling area of ​​tunnel S S=(4 / 3)ab ③The top pressure value of each meter of tunnel is Q, then: Q=Srk Where: rk—is the bulk density of roof coal rock, unit is kN / m³, take 24; ④The length L of the dangerous rock fall that the anchor cable can withstand, L = W / (S·rk·n) Where: W is the ultimate breaking force of the point anchor cable, which is 320KN; n is the safety factor, which is 1.5. BRIEF DESCRIPTION OF THE DRAWINGS

[0015] Figure 1 This is an example diagram of combined support of anchor rods and cables; Figure 2 This is an example diagram of full anchor cable support. DETAILED DESCRIPTION

[0016] In order to make the purpose, technical scheme and advantages of the embodiments of the present invention clearer, the technical scheme in the embodiments of the present invention will be clearly and completely described below. Obviously, the described embodiments are only part of the embodiments of the present invention, not all of the embodiments. The following description of at least one exemplary embodiment is actually only illustrative and is by no means intended to limit the present invention and its application or use. Based on the embodiments of the present invention, all other embodiments obtained by ordinary technicians in this field without creative work are within the scope of protection of the present invention.

[0017] The above-mentioned purpose of the present invention is achieved through the following technical solutions.

[0018] The present invention is an asymmetric support method for a large-angle fully-mechanized caving working face drift, which solves the problems of coal seam collapse and low production efficiency caused by asymmetric phenomena by using different support methods according to the stability of the drift.

[0019] First, it is necessary to calculate the safe span of rock stratum stability. When determining the safe span of rock stratum stability, it is also necessary to consider the impact of mining, burial depth, ground stress, rock strength, rock mass integrity, rock stratum collapse angle and other factors on the stability of the roof. The specific formula is as follows: Generally, when the thickness of the immediate roof is 4 times the mining height, the mining concentration coefficient is the smallest. When the tunnel is affected by mining once, the mining concentration coefficient is taken as 2.5, and the influence coefficient when affected by mining once is determined as: Determine the influence coefficient of burial depth as: For the same mine, the distribution of ground stress is basically the same, and the influence coefficient of ground stress on tunnel stability is: The rock mass strength is very different from the rock strength measured in the laboratory. Therefore, in the process of roof stability classification, it is necessary to consider the strength difference between rock and rock mass and determine that the rock mass strength is 0.65 times the rock strength, that is, The influence coefficient of rock formation integrity is determined as: Where: —Thickness of rock formations; —Effective thickness of rock formation.

[0020] Finally, the safe span for the rock formation to maintain its stability is obtained: Furthermore, the tunnel span is calculated. When calculating, firstly, according to the on-site geological conditions and the principle of ensuring safety in production, the appropriate rock layer collapse angle is selected, and then the formula is used to calculate the stable span of the roof rock layer. for: Where: —tunnel span, m; —Stable distance from the top plate, m.

[0021] —The collapse angle of the tunnel.

[0022] Furthermore, considering the main factors affecting the stability of the roof, the distance between the hard rock layer and the roof , stable span of hard rock formations and roadway span The roof stability classification index of the drift is determined as follows: Class II: Stable, m has stable rock formations, ; Class III: moderately stable, m has stable rock formations, ; Category IV: Unstable, m has stable rock formations, ; Class V: Extremely unstable, m has stable rock formations, .

[0023] Furthermore, by actually measuring the mine data and comparing it with the above-mentioned classification indicators of stability, the support method of the drift is obtained.

[0024] Furthermore, for moderately stable or unstable situations, a combined anchor rod and anchor cable support method is selected, and for extremely unstable situations, a full anchor cable support method is selected.

[0025] Furthermore, the required anchor support parameters are calculated using the formula as follows: (1) Determination of anchor rod length L=L1+L2+L3 Where: L—anchor length, m; L1—anchor exposed length, m; L2—anchor effective length, m; L3—anchor anchor length, m. Determination of anchor exposed length L1: L1 = pad thickness + nut thickness + (0.02 ~ 0.03) m Generally L1=0.15m Determination of the effective length L2 of the anchor rod: There are three methods to determine the effective length of the anchor rod. One is to use the sonic method to measure the loosening range of the tunnel surrounding rock; the second is to use a rock formation detection analyzer for measurement; the third is to use the Pugh natural equilibrium arch theory to determine. The effective length L2 of the tunnel top anchor rod is calculated as follows: The effective length L2 of the tunnel side anchor is calculated as follows: Where: f is the Proctor coefficient; B is the tunnel span, m; H is the tunnel excavation height, m; φ is the internal friction angle, °.

[0026] (2) Determination of spacing between anchor rods When supporting the roadway with anchor rods, the construction technology is usually taken into consideration and the spacing between rows is equal. The spacing between anchor rods D is calculated as follows: D≤0.5L~0.6L (m) For tunnel support, considering the influence of advance support pressure, the spacing between anchor rods can be appropriately reduced.

[0027] (3) Determination of anchor rod diameter The anchor diameter d is usually calculated as follows: In order to increase the tunnel support strength, the anchor rod diameter can be appropriately increased.

[0028] (4) Determination of anchoring force of anchor bolt Anchoring force of top anchor: Q = 3 × 1.32 × 1 × 1.5 = 5.94t Anchoring force of anchor rod: Q = 3 × 1.05 × 1 × 1.2 = 3.78 t The anchoring force of the top anchor rod is Q≥5.94t, and the anchoring force of the auxiliary anchor rod is Q≥3.78t.

[0029] Furthermore, the anchor cable support parameters are calculated using the formula. The specific calculation formula is as follows: (1) Anchor cable length According to the design principle of anchor cable support, the length of the anchor cable depends on the rock properties of the tunnel roof, rock structure, tunnel engineering dimensions and the position of the stable rock layer. The anchor cable anchoring section should be designed within 1.0m in the stable rock layer.

[0030] According to the lithology analysis of the working face, the anchor cable L 锚索 =L 外露 +L 顶煤 +L 稳定岩层 In the formula: L anchor cable - anchor cable length, unit: m; L exposed - exposed length of anchor cable (generally 0.25m); L top coal - arranged along the roof of the coal seam; L hard rock - length of the anchor cable anchoring section in the stable rock layer, taken as 1.0m.

[0031] (2) Spacing between anchor cables The spacing between anchor cables is determined based on the dangerous rock load with potential collapse tendency that the anchor cables need to bear when the anchor rod fails.

[0032] ① Natural equilibrium height b of the tunnel.

[0033] b=a / fk In the formula: a is half of the tunnel excavation width, unit: m; fk is the Proctor coefficient of the roof rock.

[0034] ② Potential falling area of ​​tunnel S S=(4 / 3)ab ③The top pressure value of each meter of tunnel is Q, then: Q=Srk Where: rk is the bulk density of roof coal rock, unit is kN / m³, take 24.

[0035] ④The length L of the dangerous rock fall that the anchor cable can withstand, L = W / (S·rk·n) Where: W is the ultimate breaking force of the point anchor cable, which is 320KN; n is the safety factor, which is 1.5.

[0036] Furthermore, the above calculations determine the spacing and length of anchor cables and anchor rods when two support forms are used. The number of anchor rods and anchor cables to be installed is determined by the spacing of anchor rods and anchor cables, the length of the trench, and the actual situation.

[0037] Furthermore, when the combined support method of anchor rods and anchor cables is adopted, about six long anchor rods are set on the top plate, about five are arranged on each side, and 3 / 4 anchor cables are arranged on the top. The anchor rods and anchor cables are staggered with a step distance of about 800mm.

[0038] Furthermore, when full anchor cable support is adopted, about seven anchor cables are arranged at the top, about five cables on each side, and the pitch is about 800mm. During actual placement, adjustments can be made according to the above calculation formula and actual conditions.

[0039] Furthermore, it should be noted that when carrying out actual support, the length and number of anchor rods and cables should be increased or decreased accordingly according to the asymmetric location to achieve a higher effect.

[0040] Through the above methods, the stability of the support can be appropriately increased, the coal seam collapse and other problems can be prevented, the deformation of the coal seam roof can be greatly reduced, and the efficient generation of the working face can be ensured.

Claims

1. A method for asymmetric support of a large-angle fully-mechanized caving working face, characterized in that: include: S1: Calculate the safe span of rock stability, including mining influence coefficient, burial depth influence coefficient, ground stress influence coefficient, rock mass strength influence, and rock integrity influence; S2: Calculate the tunnel span; S3: According to the distance S between the hard rock layer and the roof, the stable span L of the hard rock layer and the change of the tunnel span B, the classification index of the stability of the coal mine drift roof is determined; S4: According to the actual data measurement of the coal mine, the basic form and main parameters of the drift support are determined; S5: Divide the drift into areas according to the stability of the roof, and support them in sections according to the corresponding support methods. When encountering a roof with poor stability, choose an asymmetric support method for support; S6: For moderately stable or unstable roofs, the asymmetric support of the trench adopts the combined support of anchor rods and cables; for extremely unstable roofs, the asymmetric support of the trench adopts the full anchor cable support; S7: Calculate the parameters related to anchor support; S8: Calculate the parameters related to anchor cable support; S9: Select the appropriate number of anchor rods and cables to provide corresponding asymmetric support for the drift.

2. The asymmetric support method for the large-angle fully-mechanized caving working face according to claim 1 is characterized in that: The step S1 comprises: When the thickness of the immediate roof is 4 times the mining height, the mining concentration coefficient is the smallest. When the tunnel is affected by mining once, the mining concentration coefficient is taken as 2.5, and the influence coefficient when affected by mining once is determined as: Determine the influence coefficient of burial depth as: For the same mine, the distribution of ground stress is basically the same, and the influence coefficient of ground stress on tunnel stability is: The rock mass strength is very different from the rock strength measured in the laboratory. Therefore, in the process of roof stability classification, it is necessary to consider the strength difference between rock and rock mass and determine that the rock mass strength is 0.65 times the rock strength, that is, The influence coefficient of rock formation integrity is determined as: Where: —Thickness of rock formations; —Effective thickness of rock formation; Finally, the safe span for the rock formation to maintain its stability is obtained: 。 3. The asymmetric support method for the large-angle fully-mechanized caving working face according to claim 1 is characterized in that: In step S2, the stable span B of the roof rock layer is: Where: —tunnel span, m; —Stable distance from the top plate, m. —The collapse angle of the tunnel.

4. The asymmetric support method for the drift of a large-angle fully-mechanized caving working face according to claim 1 is characterized in that: In step S3, considering the main influencing factors of roof stability, according to the distance s between the hard rock layer and the roof, the stable span L of the hard rock layer and the change of the tunnel span B, the roof stability classification index of the drift is determined as follows: Class II: Stable, m has stable rock formations, ; Class III: moderately stable, m has stable rock formations, ; Category IV: Unstable, m has stable rock formations, ; Class V: Extremely unstable, m has stable rock formations, .

5. The asymmetric support method for the drift of a large-angle fully-mechanized caving working face according to claim 1 is characterized in that: In step S7, the required anchor support parameters are calculated using the following formula: (1) Determination of anchor rod length L=L1+L2+L3 Where: L—anchor length, m; L1—anchor exposed length, m; L2—anchor effective length, m; L3—anchor anchor length, m. Determination of anchor exposed length L1: L1 = pad thickness + nut thickness + (0.02 ~ 0.03) m Determination of the effective length L2 of the anchor rod: There are three methods to determine the effective length of the anchor rod. One is to use the sonic method to measure the loosening range of the surrounding rock of the tunnel; the second is to use a rock formation detection analyzer for measurement; the third is to use the Pugh natural equilibrium arch theory to determine it; The effective length L2 of the tunnel top anchor is calculated as follows: The effective length L2 of the tunnel side anchor is calculated as follows: Where: f—Praeter coefficient; B—tunnel span, m; H—tunnel excavation height, m; φ—internal friction angle, °; (2) Determination of spacing between anchor rods When supporting the roadway with anchor rods, the construction technology is usually taken into consideration and the spacing between rows is equal. The spacing between anchor rods D is calculated as follows: D≤0.5L~0.6L For tunnel support, considering the influence of advance support pressure, the spacing between anchor bolts can be appropriately reduced; (3) Determination of anchor rod diameter The anchor diameter d is usually calculated as follows: In order to increase the tunnel support strength, the anchor diameter can be appropriately increased; (4) Determination of anchoring force of anchor bolt Anchoring force of top anchor: Q = 3 × 1.32 × 1 × 1.5 = 5.94t Anchoring force of anchor rod: Q = 3 × 1.05 × 1 × 1.2 = 3.78 t The anchoring force of the top anchor rod is Q≥5.94t, and the anchoring force of the auxiliary anchor rod is Q≥3.78t.

6. The asymmetric support method for the drift of a large-angle fully-mechanized caving working face according to claim 1 is characterized in that: In step S8, the anchor cable support parameters are calculated using a formula, and the specific calculation formula is as follows: (1) Anchor cable length According to the design principle of anchor cable support, the length of the anchor cable depends on the lithology of the roadway roof, the rock mass structure, the size of the roadway project and the position of the stable rock layer. The anchor cable anchoring section should be designed within 1.0m in the stable rock layer. According to the lithology analysis of the working face, the anchor cable L 锚索 =L 外露 +L 顶煤 +L 稳定岩层 Where: L anchor cable - anchor cable length, unit: m; L exposed - exposed length of anchor cable; L top coal - arranged along the roof of coal seam; L hard rock - length of stable rock layer of anchor cable anchoring section, take 1.0m; (2) Spacing between anchor cables The spacing between anchor cables is determined based on the dangerous rock load with potential collapse tendency that the anchor cables need to bear when the anchor rod fails. ① Natural equilibrium height b of the tunnel; b=a / fk Where: a is half of the tunnel excavation width, in m; fk is the Proctor coefficient of the roof rock; ② Potential falling area of ​​tunnel S S=(4 / 3)ab ③The top pressure value of each meter of tunnel is Q, then: Q=Srk Where: rk—is the bulk density of roof coal rock, unit is kN / m³, take 24; ④The length L of the dangerous rock fall that the anchor cable can withstand, L = W / (S·rk·n) Where: W is the ultimate breaking force of the point anchor cable, which is 320KN; n is the safety factor, which is 1.5.

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