CN111022048A - Fully mechanized coal mining face withdrawing and final mining method - Google Patents

Fully mechanized coal mining face withdrawing and final mining method Download PDF

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CN111022048A
CN111022048A CN201911122862.0A CN201911122862A CN111022048A CN 111022048 A CN111022048 A CN 111022048A CN 201911122862 A CN201911122862 A CN 201911122862A CN 111022048 A CN111022048 A CN 111022048A
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anchor
roof
mining
top plate
withdrawing
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CN111022048B (en
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谷拴成
孙魏
王盼
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Xian University of Science and Technology
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Xian University of Science and Technology
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    • E—FIXED CONSTRUCTIONS
    • E21—EARTH DRILLING; MINING
    • E21C—MINING OR QUARRYING
    • E21C41/00—Methods of underground or surface mining; Layouts therefor
    • E21C41/16—Methods of underground mining; Layouts therefor
    • E21C41/18—Methods of underground mining; Layouts therefor for brown or hard coal
    • E—FIXED CONSTRUCTIONS
    • E21—EARTH DRILLING; MINING
    • E21D—SHAFTS; TUNNELS; GALLERIES; LARGE UNDERGROUND CHAMBERS
    • E21D20/00—Setting anchoring-bolts

Abstract

The invention discloses a fully mechanized coal mining face withdrawing and final mining method, which comprises the following steps: installing a bracket in the retraction channel; adjusting the bottom plate of the working surface to be 180-220 mm higher than the bottom plate of the withdrawing channel; determining support of an anchor rod and an anchor cable of the withdrawal channel, and then supporting the withdrawal channel by using the anchor rod and the anchor cable; carrying out end mining and net hanging coal mining; stopping mining and withdrawing the fully mechanized mining equipment. The method of the invention is based on the support determination method of the roof anchor rod, the roof anchor cable and the side anchor rod to carry out final mining, and can avoid support reworking and ensure that the final mining is carried out smoothly on the basis of once completing the support design.

Description

Fully mechanized coal mining face withdrawing and final mining method
Technical Field
The invention belongs to the technical field of mine mining, and particularly relates to a fully mechanized coal mining face withdrawing and final mining method.
Background
The fully mechanized coal mining face withdrawal directly influences the coal mine production efficiency, and safety, high efficiency and economy are three important indexes. The stability of the surrounding rock of the withdrawal passage is the premise of ensuring the safety and the rapid withdrawal of equipment, the production efficiency and the economic benefit of a coal mine are directly concerned, and coal mine enterprises and experts in the industry pay attention to the problems. Under the dual action of the excavation and mining influences of the withdrawal roadway, the property of surrounding rock of the roadway can change, so that the structure and the motion rule of a coal seam roof of a mining area are fully considered when the withdrawal roadway support design is carried out, and the support design is carried out on the withdrawal roadway.
According to the position of a withdrawal roadway, the condition of an underground rock stratum and the development condition of rock stratum fractures, the support design of the current withdrawal roadway faces two difficulties: firstly, in the working face propelling process, the change rule of the stress of surrounding rocks of the pre-digging and withdrawing roadway is complex, and the deformation of the surrounding rocks is large; secondly, the fracture and the self mechanical property of the roof rock stratum of the withdrawn roadway are not easy to determine.
Poor support forms can lead to the support being pressed down, brings danger for the work of moving, reduces production efficiency, and too conservative support forms also can bring inconvenience for moving, reduces production efficiency. Therefore, when the withdrawal roadway is designed, not only the influence of working face mining on roadway surrounding rocks is considered fully, but also the mechanical property of the roadway roof rock stratum is considered, the support is reduced as much as possible under the condition that potential safety hazards are not brought, and the self bearing capacity of the roadway roof rock stratum is brought into full play. Aiming at the problems, it is necessary to provide a feasible fully mechanized coal mining face withdrawing method.
Disclosure of Invention
The invention aims to solve the technical problem of providing a fully mechanized coal mining face withdrawing and final mining method aiming at the defects of the prior art. The method can avoid the reworking of the support on the basis of completing the support design once, and ensures that the final mining is carried out smoothly.
In order to solve the technical problems, the invention adopts the technical scheme that: a fully mechanized coal mining face withdrawing and final mining method is characterized by comprising the following steps:
installing a bracket in the retraction channel;
adjusting the bottom plate of the working surface to be 180-220 mm higher than the bottom plate of the withdrawing channel;
determining support of an anchor rod and an anchor cable of the withdrawal channel, and then supporting the withdrawal channel by using the anchor rod and the anchor cable;
carrying out end mining and net hanging coal mining;
stopping mining and withdrawing the fully mechanized mining equipment.
The fully mechanized coal mining face withdrawing end mining method is characterized by further comprising analyzing and predicting the periodic pressure of a withdrawing channel; the analyzing and predicting the periodic back pressure of the withdrawal passage comprises the following steps:
when the working face is recovered to 200m, carrying out first analysis and prediction on the periodic incoming pressure rule;
when the working face is recovered to 200-100 m, analyzing and predicting the periodic pressure-incoming rule every 50 m;
when the working face is recovered to be between 100m and 50m, the periodic pressure-incoming rule is analyzed and predicted every 25 m.
The fully mechanized coal mining face withdrawing and final mining method is characterized in that when the result of the first analysis and prediction is that the old roof is broken at an unfavorable position when the old roof is through, the mining height or the mining speed needs to be adjusted, and the result after the first measure is taken is evaluated when the analysis and prediction is carried out for the second time; the unfavorable position is right above the bracket; the support comprises a stack support or a hydraulic support.
The fully mechanized coal mining face withdrawing and final mining method is characterized in that the mounting stack type support is arranged between the working face withdrawing distance and the main withdrawing channel distance of 150-100 m.
The fully mechanized coal mining face withdrawing and final mining method is characterized in that the adjusting working face bottom plate is withdrawn on the working face to a distance of 20-30 m from the main withdrawing channel.
The fully mechanized mining face withdrawing and final mining method is characterized in that the method for determining the support of the anchor rod and the anchor cable of the withdrawing channel comprises the following steps:
step one, calculating the radius of a top plate fracture area and the radius of a top plate plastic area;
secondly, calculating parameters of the roof anchor rods, wherein the parameters of the roof anchor rods comprise the length of the roof anchor rods, the row spacing between the roof anchor rods and the prestress of the roof anchor rods;
calculating parameters of the top plate anchor cables, wherein the parameters of the top plate anchor cables comprise the length of the top plate anchor cables, the row distance between the top plate anchor cables and the prestress of the top plate anchor cables;
and step four, calculating parameters of the upper anchor rods, wherein the parameters of the upper anchor rods comprise the length of the upper anchor rods, the row spacing among the upper anchor rods and the prestress of the upper anchor rods.
The fully mechanized mining face withdrawal final mining method is characterized in that the calculation method of the radius of the roof fracture area and the radius of the roof plastic area in the step one is calculated according to the following formula:
radius R of top plate crushing zonebThe calculation formula of (2) is as follows:
Figure RE-GDA0002385873100000031
the calculation formula of the radius of the plastic area of the top plate is as follows:
Figure RE-GDA0002385873100000032
wherein the content of the first and second substances,
Figure RE-GDA0002385873100000033
Figure RE-GDA0002385873100000034
Figure RE-GDA0002385873100000035
Figure RE-GDA0002385873100000036
Figure RE-GDA0002385873100000037
σcultimate compressive strength in MPa;
Figure RE-GDA0002385873100000038
the compressive residual strength of the rock mass is in MPa;
r0is the equivalent circle radius of the withdrawal passage, and the unit is m;
q is the softening modulus of the strength of the roadway rock mass, and the unit is GPa;
qithe strength of the collapse zone is MPa;
m is a correlation coefficient of the expansion fragmentation property of the surrounding rock plastic zone;
k is a plasticity coefficient;
P0is the stress of the original rock and has the unit of MPa;
mu is Poisson's ratio;
σRis the radial stress of the elastoplastic interface, in MPa;
e is the modulus of elasticity in GPa.
The fully mechanized mining face withdrawing and final mining method is characterized in that in the second step, the calculation method of the parameters of the roof bolt is calculated according to the following formula:
roof anchor anchoring length LaThe calculation formula of (2) is as follows:
Figure RE-GDA0002385873100000041
wherein the content of the first and second substances,
k1the roof anchor rod anchoring force safety coefficient is set;
P1the bearing capacity of the roof anchor rod is kN;
d is the diameter of the anchor rod hole of the top plate, and the unit is mm;
τrthe unit is the bonding strength between the anchoring agent and the rock mass and is MPa;
the calculation formula of the roof bolt length L is as follows:
L＝La+a1+Lb； (4)
wherein the content of the first and second substances,
Lathe anchoring length of the roof anchor rod is m;
Lbthe exposed length of the roof bolt is m;
a1the depth of a crushing area of a top plate of the withdrawal channel is m;
the calculation formula of the row spacing a between the roof anchor rods is as follows:
Figure RE-GDA0002385873100000042
wherein, γ1Is the coal bed severity with the unit of kN/m3；
G is the anchoring force of the roof anchor rod, and the unit is kN;
s1the safety factor of the row spacing between the roof anchor rods is set;
the calculation formula of the roof bolt prestress N is as follows:
Figure RE-GDA0002385873100000051
wherein gamma is the average top plate weight and has a unit of kN/m3；
n is the distribution density of roof bolts in root/m2；
Delta is the thickness of the lower rock stratum and is expressed in m;
α is the formation dip angle in °;
fais the interlayer friction coefficient.
The fully mechanized mining face withdrawal and final mining method is characterized in that in the third step, the calculation method of the parameters of the roof anchor cable is calculated according to the following formula:
anchoring length b of anchor cable on top plate2The calculation formula of (2) is as follows:
Figure RE-GDA0002385873100000052
wherein the content of the first and second substances,
K1the safety coefficient of the anchorage force of the anchor cable of the top plate is set;
P2the bearing capacity of the anchor cable of the top plate is kN;
d is the diameter of the anchor cable hole drilling hole of the top plate, and the unit is mm;
the calculation formula of the arch height b of the natural balance arch is as follows:
Figure RE-GDA0002385873100000053
wherein the content of the first and second substances,
b is the span of the withdrawal passage, and the unit is m;
a2the plastic zone depth of the upper part is m;
fbis the coefficient of Pyth; the length L of the roof anchor cabledThe calculation formula of (2) is as follows:
Ld＝b+b1+b2(9)
wherein the content of the first and second substances,
b is the arch height of the natural balance arch, and the unit is m;
b1the exposed length of the anchor cable of the top plate is m;
b2the anchoring length of the anchor cable of the top plate is m;
the calculation formula of the row spacing between the roof anchor cables is as follows:
Figure RE-GDA0002385873100000054
wherein the content of the first and second substances,
PNbreaking force is provided for each row of anchor cables; the unit is kN;
K3the safety factor of the row spacing between the anchor cables is set;
the roof anchor cable prestress N3The calculation formula of (2) is as follows:
Figure RE-GDA0002385873100000061
n3the distribution density of the anchor cables of the top plate is in the unit of root/m2。
The fully mechanized mining face withdrawing and final mining method is characterized in that the calculation method of the parameters of the side anchor rod in the fourth step is calculated according to the following formula:
anchoring length L of anchor rod on side2The calculation formula of (2) is as follows:
Figure RE-GDA0002385873100000062
k2the safety coefficient of the anchoring force of the anchor rod at the upper part is set;
P4the bearing capacity of the side anchor rod is kN;
D4the diameter of a hole drilled for an anchor rod hole of the upper part is in mm;
the side anchor rod L1The length is calculated as:
L1＝L2+a4+L4(13)
wherein the content of the first and second substances,
L2the anchoring length of the anchor rod at the upper part is m;
L4the exposed length of the anchor rod at the upper part is m;
the calculation formula of the row spacing between the side anchor rods is as follows:
Figure RE-GDA0002385873100000063
wherein the content of the first and second substances,
s2the safety factor of the row spacing between the anchor rods at the upper part is set;
the calculation formula of the prestress of the upper anchor rod is as follows:
Figure RE-GDA0002385873100000064
wherein the content of the first and second substances,
n4the distribution density of the anchor rods at the upper part is in root/m2。
Compared with the prior art, the invention has the following advantages:
1. the fully mechanized coal mining face withdrawing and final mining method is based on a support determination method of the roof anchor rod, the roof anchor cable and the side anchor rod to perform final mining, and can avoid support reworking on the basis of once completed support design and ensure that the final mining is performed smoothly.
2. The fully mechanized coal mining face withdrawing and final mining method can reduce vertical deformation of surrounding rocks of the roadway, effectively control the deformation range of the surrounding rocks, weaken the influence of mining on two withdrawing channels, effectively control the horizontal displacement of the surrounding rocks and maintain the stability of the roadway.
3. The fully mechanized coal mining face withdrawal final mining method disclosed by the invention adopts a method of combining a tunnel elastoplasticity theory and a loose ring design concept under a static pressure condition, fully considers a strain softening condition, meets the requirement of reasonable thickness of a plastic ring in coal rock engineering support, realizes the deformation energy of surrounding rock to the maximum extent, improves the bearing capacity of anchor rod and anchor cable support, and has a potential application prospect.
4. According to the method, elastic-plastic mechanical analysis is carried out on surrounding rock of the withdrawal passage model, meanwhile, the surrounding rock load is determined through a loosening zone theory, the influence range after the roadway is excavated is mainly concentrated in the loosening range of the plastic zone, and the roadway supporting load is mainly the self weight of the surrounding rock in the loosening zone.
The technical solution of the present invention is further described in detail with reference to the accompanying drawings and embodiments.
Drawings
FIG. 1 is a simplified model of tunnel elastoplasticity under static pressure.
Fig. 2 is a withdrawal passage equivalent circle.
Fig. 3 is a cross-sectional view of the original support of the N1114 working face retraction channel.
FIG. 4 shows the variation of the measuring points 2m inside the coal mining slope along with the mining stress.
FIG. 5 shows the variation of the measuring points 2m inside the non-mining slope coal body along with the mining stress.
Figure 6 shows the main retrace path roof as a function of mining vertical displacement.
Figure 7 shows the main withdrawal passage highwall as a function of production level displacement.
FIG. 8 is a vertical displacement variation curve of the lower top plate of the original support and the support scheme determined by the method of the invention.
FIG. 9 is the horizontal displacement curve of the two sides of the lower top plate of the original support and the support scheme determined by the method of the invention.
Description of the reference numerals
1-roof anchor cable 2-roof anchor rod; 3-anchor rod of upper part;
4-a steel belt; 5-lead wire mesh.
Detailed Description
The fully mechanized mining face withdrawing end mining method uses caragana microphylla coal mine 1-2And N1114 working face.
Step one, when the working face is stoped to a distance of 200m from a withdrawal passage, the working face hydraulic support is overhauled to ensure that the working face hydraulic support meets the requirement of initial supporting force of mine pressure control at the end mining stage;
secondly, when the working face is recovered to a distance of 150-100 m from the withdrawal channel, a stack type support is installed in the withdrawal channel;
thirdly, when the working face is recovered to a distance of 50m from the withdrawal channel, the hydraulic support in the working face and the stack type support in the withdrawal channel are overhauled and replenished with liquid, so that the hydraulic support of the working face meets the requirement of designing the initial supporting force;
step four, when the working face is recovered to a distance of 25m from the withdrawal channel, adjusting the bottom plate of the working face to be 180-220 mm higher than the bottom plate of the withdrawal channel;
in the first step to the fourth step, analysis and prediction are carried out on the periodic back pressure of the withdrawal passage; the analyzing and predicting the periodic back pressure of the withdrawal passage comprises the following steps:
when the working face is recovered to 200m, carrying out first analysis and prediction on the periodic incoming pressure rule; when the result of the first analysis and prediction is that the old jack is broken at an unfavorable position when the old jack is penetrated, the mining height or the mining speed is required to be adjusted, and the result after the first measure is taken is evaluated when the analysis and prediction is carried out for the second time; the unfavorable position is right above the support, and the support is a stack support or a hydraulic support;
when the working face is recovered to 200-100 m, analyzing and predicting the periodic pressure-incoming rule every 50 m;
when the working face is recovered to be between 100m and 50m, the periodic pressure-incoming rule is analyzed and predicted every 25 m.
Fifthly, when the working face is recovered to a position 20m away from the retraction channel, steel wire ropes are used for hanging the steel bar net in the auxiliary retraction channel, and standby materials are prepared; the standby material comprises a metal mesh, a scaffold, a 14# iron wire, a net hook, a hoop, a winch and a pulley;
sixthly, when the working face is pushed to the day before the stoping line, the standby material is conveyed to the position of the return air crossheading stoping line by a shovel car;
seventhly, mounting a first fixed pulley on a top beam of the hydraulic support on the working surface, mounting a winch, a hoop and a second fixed pulley on the upright column on one surface of the hydraulic support close to the withdrawal channel, and mounting steel wire ropes on the first fixed pulley and the second fixed pulley;
step eight, carrying out maintenance before final mining;
step nine, liquid is replenished into the working surface hydraulic support and the stack type support in the withdrawing channel, the frame shapes of the working surface hydraulic support and the stack type support in the withdrawing channel are adjusted, and a side protection plate of the hydraulic support and a side protection plate of the stack type support are beaten to support the coal wall;
step ten, determining support of the anchor rod and the anchor cable of the withdrawal channel, and then supporting the withdrawal channel by using the anchor rod and the anchor cable; the method for determining the support of the roof anchor rod, the roof anchor cable and the side anchor rod comprises the following steps:
according to the through-hole data and the underground measured data, the thickness of the coal seam in the N1114 working face tunneling range is 0.6-1.92 m, the average thickness is 1.75m, the coal seam inclination angle is 0 degree 41 '-1 degree 55', and the coal seam locally fluctuates; the coal seam burial depth is 50-150 m, wherein the thickness of the bedrock is 45-55 m, and the thickness of the soil layer is 10-100 m. 1-2The coal layer is semi-dark molded coal, the silky luster is glossy, the upper part of the coal layer contains a layer of gangue, the thickness of the gangue is 0.1-0.2 m, and the gangue is carbonaceous mudstone and siltstone.
The pseudo top of the coal bed is a dark gray mudstone coal clamping line with the thickness of 0m to 0.39 m; the direct top is gray, dark gray siltstone and silty mudstone, or light gray-gray siltstone and fine sandstone which are laminated, the thickness is 0-1.51 m, the old top is fine sandstone, the thickness is 3.43-12.16 m, and the local slow wave-shaped layer is formed. The bottom plate is made of gray or dark gray siltstone or mudstone, and the local part is made of black carbon mudstone with the thickness of 0.37-4.92 m.
N1114 working face retract borehole geological data near the channel: the span of the withdrawal passage is 5.3 m; height 2.65 m; the thickness of the covering layer in the region is 50-150 m, and the maximum value of the buried depth in the calculation process is 150 m; the average thickness of the coal bed is 1.75 m; in the coal seamAngle of friction
Figure RE-GDA0002385873100000091
The cohesive force C of the coal seam is 0.73 MPa; coal bed severe gamma1＝12kN/m3(ii) a The equivalent circle radius r of the withdrawal passage is 2.96m, and the average gravity gamma of the top plate is 24.5kN/m3The cohesion C of the top plate is 1.86MPa, and the internal friction angle is
Figure RE-GDA0002385873100000092
Under the static pressure condition, the elastoplasticity of the round tunnel simplifies the model, the radius of the tunnel is r0The broken area and plastic area of the top plate are shown in figure 1.
Step 101, calculating the radius of a top plate fracture area and a plastic area; according to a method for combining a tunnel elastoplasticity theory and a loosening zone theory under a static pressure condition, determining the load of surrounding rocks by the loosening zone theory, wherein the influence range of the excavated tunnel is mainly concentrated in the loosening range of a plastic zone, and the supporting load of the tunnel is mainly the self weight of the surrounding rocks in the loosening zone;
the calculation formula of the radius of the top plate fracture area and the top plate plastic area is as follows:
Figure RE-GDA0002385873100000101
Figure RE-GDA0002385873100000102
Figure RE-GDA0002385873100000103
Figure RE-GDA0002385873100000104
Figure RE-GDA0002385873100000105
Figure RE-GDA0002385873100000106
Figure RE-GDA0002385873100000107
Figure RE-GDA0002385873100000108
Figure RE-GDA0002385873100000109
wherein the content of the first and second substances,
σc7.645MPa for ultimate compressive strength;
Figure RE-GDA00023858731000001011
the compressive residual strength of the rock mass is 2.13 MPa;
the ultimate compressive strength and the compressive residual strength of the rock mass can be tested by a rigidity testing machine to obtain data;
r02.96m for the equivalent circle radius of the withdrawal passage;
the calculation method of the equivalent circle radius of the withdrawal passage comprises the following steps: as shown in fig. 1, the span of the rectangular retrace channel is 2a, the height of the rectangular channel is 2b, the rectangular retrace channel is equivalent to a circle, as shown in fig. 2, according to the formula
Figure RE-GDA00023858731000001010
Calculating to obtain the equivalent circle radius r of the withdrawal passage0；
Q is the softening modulus of the strength of the roadway rock mass, and is 4.7 GPa;
Figure RE-GDA0002385873100000111
the critical tangential strain of the rock mass entering the plastic residual stage from the plastic softening stage can be obtained by testing through a rigidity testing machine;
Figure RE-GDA0002385873100000112
the critical tangential strain of the rock body when the rock body transits from the elastic deformation stage to the plastic softening stage is 5.12 multiplied by 10 data obtained by the test of a rigidity tester-3；
qiThe intensity of the caving zone is equal to the average supporting stress of the top plate, equal to the total support working resistance/(the length of the roadway multiplied by the width of the roadway), and is 0.38 MPa;
m is the correlation coefficient of the expansion fragmentation property of the plastic zone of the surrounding rock,
Figure RE-GDA0002385873100000113
Figure RE-GDA0002385873100000114
the internal friction angle is 38.11 degrees;
k is the coefficient of plasticity of the mixture,
Figure RE-GDA0002385873100000115
P0the stress of the original rock is 9.6 MPa;
mu is Poisson's ratio, 0.33;
σRis the radial stress of the elastoplastic interface, MPa,
Figure RE-GDA0002385873100000116
e is the elastic modulus, 10.186 GPa;
to obtain Rb＝3.23m；Rp＝4.03m；
The crushing zone depth a of the top plate of the withdrawal passage1＝Rb-withdrawal channel height/2-3.23-2.65/2-1.91 m;
depth a of crushing zone of upper part of withdrawal channel2＝Rb-withdrawal channel span/2-3.23-5.3/2-0.58 m;
plastic region depth a of top plate of withdrawal passage3＝Rp-withdrawal channel height/2-4.03-2.65/2-2.705 m;
plastic zone depth a of upper part of withdrawal channel4＝Rp-withdrawal channel span/2 ═ 4.03-5.3/2＝1.38m。
102, calculating parameters of the roof anchor rods 2, wherein the parameters of the roof anchor rods comprise the length of the roof anchor rods, the row spacing between the roof anchor rods and the prestress of the roof anchor rods;
the roof bolt adopts a phi 20 left-handed thread steel bolt, and the bearing capacity P is 105 kN;
anchoring length L of roof boltaComprises the following steps:
Figure RE-GDA0002385873100000117
wherein the content of the first and second substances,
k1the roof anchor rod anchoring force safety factor is 1;
P1the bearing capacity of the roof bolt is 105 kN;
d is the diameter of a drill hole of the anchor rod hole of the top plate, and is 32 mm;
τr2MPa, which is the bonding strength between the anchoring agent and the rock mass;
bring in data to obtain La＝0.52m；
The top stock length L is:
L＝La+a1+Lb(4)；
Lathe anchoring length of the roof bolt is 0.52 m;
Lbthe exposed length of the roof bolt is 0.1 m;
a11.91m for retracting the crushing zone depth of the channel top plate;
the length of the roof anchor rod is 2.53 m;
the row spacing a between the roof anchor rods is as follows:
Figure RE-GDA0002385873100000121
wherein, γ1The coal bed severity is obtained; 12kN/m3(ii) a The coal bed severity is the natural severity of the coal bed and is the ratio of the coal weight to the coal volume;
Lathe anchoring length of the roof bolt is 0.52 m;
g is the anchoring force of the anchor rod, and is 50 kN;
s1the safe coefficient of the row spacing between the roof anchor rods is 4;
the row spacing between roof bolts is 1.42 m.
The prestress of the roof bolt is as follows:
Figure RE-GDA0002385873100000122
gamma is the average overburden gravity of 24.5kN/m3；
n is the distribution density of the roof bolt, 0.8 bolts/m2；
Delta is the thickness of a lower rock stratum, and the thickness of the lower rock stratum is the sum of the thickness of a direct roof and the thickness of a coal seam pseudo roof; the thickness of the lower rock stratum is different according to the stratum condition, the surrounding rock property and the mining condition; δ 1.51+0.39 1.9 m;
α is the formation dip, 1 °;
fais the interlayer friction coefficient, 1; the interlayer friction coefficient refers to the ratio of the friction force between rock layers to the vertical force vertical to the friction surface;
the prestress N of the roof anchor rod is not less than 59.1kN, and the minimum prestress of the roof anchor rod is 60 kN;
and (3) determining the supporting parameter scheme of the roof bolt of the main withdrawing channel by combining the technical requirements on the bolt supporting design of the GB35056-2018 coal mine roadway bolt supporting technical specification as follows: phi 20 multiplied by 2600mm left-handed non-longitudinal rib deformed steel bar anchor rods are arranged at intervals of 800mm multiplied by 1000mm, wherein the interval is the interval multiplied by the row pitch; the installation angle of the roof anchor rod close to the roadway side is 10 degrees to the vertical line, the anchoring length is 600mm, and the pretightening force is applied to 60 kN.
Step 103, calculating parameters of the top plate anchor cable 1, wherein the parameters specifically comprise the length of the top plate anchor cable, the row distance between the top plate anchor cables and the prestress of the top plate anchor cable;
roof anchor cable is diameter
Figure RE-GDA0002385873100000133
The bearing capacity of the steel strand is 320kN, the breaking force is 355kN, and the anchor rope anchor of the top plateFixed length b2Comprises the following steps:
Figure RE-GDA0002385873100000131
wherein the content of the first and second substances,
K 11, a roof anchor cable anchoring force safety factor;
P2the bearing capacity of the anchor cable of the top plate is 320 kN;
d is the diameter of a drilled hole of the anchor cable hole of the top plate, and is 32 mm;
τr2MPa, which is the bonding strength between the anchoring agent and the rock mass;
the anchor length b of the anchor cable of the top plate is obtained2Is 1.59 m;
the free section length of the roof anchor cable is calculated according to the strain softening and natural balance arch theory, and the calculation formula of the arch height b of the natural balance arch is as follows:
Figure RE-GDA0002385873100000132
wherein the content of the first and second substances,
b is the span of the withdrawal passage, 5.3 m;
a2the upper plastic zone depth is 1.38 m;
fb1.1 is the coefficient of Pyth; the Pythian coefficient is also called the firmness coefficient or the fastening coefficient of the rock;
the arch height of the natural balance arch is 3.66 m;
the length L of the roof anchor cabledComprises the following steps:
Ld＝b+b1+b2(9)
wherein the content of the first and second substances,
b is the arch height of the natural balance arch, 3.66 m;
b1the exposed length of the anchor cable of the top plate is 0.3 m;
b2the anchoring length of the anchor cable for the top plate is 1.59 m;
the length of the anchor cable brought into the top plate is 5.55 m;
the row spacing s between the roof anchor cables is as follows:
Figure RE-GDA0002385873100000141
wherein the content of the first and second substances,
b is the span of the withdrawal passage, 5.3 m;
PNfor each row of anchor cables, the breaking force of a single anchor cable is 355kN, the number of the anchor cables in each row is 4, and the breaking force of each row of anchor cables is 355kN multiplied by 4 which is 1420 kN.
K3The safety factor of the row spacing between the anchor cables is 1.8;
γ1the coal bed is 12kN/m in severe degree3(ii) a Bringing s to 2.54 m;
the roof anchor cable prestress N3Comprises the following steps:
Figure RE-GDA0002385873100000142
gamma is the mean top plate weight, 24.5kN/m3；
n3The density of the anchor cables of the top plate is 0.3 per meter2；
Delta is the thickness of a lower rock stratum, and the thickness of the lower rock stratum is the sum of the thickness of a direct roof and the thickness of a coal seam pseudo roof; the thickness of the lower rock stratum is different according to the stratum condition, the surrounding rock property and the mining condition; taking the delta as 1.51+0.39 as 1.9 m;
α is the formation dip, 1 °;
fbis the interlayer friction coefficient, 1;
the prestress of the top plate anchor cable is 157.85kN, and the prestress of the top plate anchor cable is 160 kN.
And (3) combining the requirements on anchor rod (cable) design from the above technical specifications of rock and soil anchor rods (cables), and determining the support parameters of the anchor cable of the top plate of the main withdrawing channel as follows: the length of the anchor cable is 6000mm, the anchoring length is 1800mm, the row spacing is 2150 multiplied by 2000mm, and the prestress applied to the anchor cable is 160 kN.
104, calculating parameters of the upper anchor rods 3, wherein the parameters of the upper anchor rods comprise the length of the upper anchor rods, the row spacing among the upper anchor rods and the prestress of the upper anchor rods;
the anchor rod at the upper part is a glass fiber reinforced plastic anchor rod with phi 18, and the bearing capacity is 85 kN;
anchoring length L of anchor rod on side2Comprises the following steps:
Figure RE-GDA0002385873100000151
wherein k is2The safety factor of the anchoring force of the anchor rod at the upper part is 1;
P4the bearing capacity of the side anchor rod is 85 kN;
D4drilling a hole with the diameter of 32mm for the anchor rod hole of the upper part;
τrthe bonding strength between the anchoring agent and the slope rock mass is 2 MPa;
the anchoring length of the anchor rod brought into the upper part is 0.42 m;
length L of anchor rod1Comprises the following steps:
L1＝L2+a4+L4(13)
L2the anchoring length of the anchor rod at the upper part is 0.42 m;
L4the exposed length of the anchor rod at the upper part is 0.1 m;
a4the plastic zone depth of the channel upper part is withdrawn, and is 1.38 m;
the length of the anchor rod at the upper part is 1.9 m;
the row spacing h between the anchor rods of the upper part is as follows:
Figure RE-GDA0002385873100000152
wherein the content of the first and second substances,
γ1is the coal bed severe, 12kN/m3；
L2The anchoring length of the anchor rod at the upper part is 0.42 m;
g is the anchoring force of the anchor rod at the upper part, and is 50 kN;
s2the safety factor of the row spacing between the side anchor rods is 4;
the row spacing h between the anchor rods of the upper part is 1.57 m.
The upper part anchorRod prestress N4Comprises the following steps:
Figure RE-GDA0002385873100000153
wherein the content of the first and second substances,
γ1the coal bed is 12kN/m in severe degree3；
n4The density of the anchor rod at the upper part is 1 root/m2；
α is the angle of inclination of the formation, 1,
f is the interlayer friction coefficient, 1;
a4the plastic zone depth of the channel upper part is withdrawn, and is 1.38 m;
the prestress N of the side anchor rod is more than or equal to 16.85kN, and the minimum prestress of the side anchor rod is 20 kN;
and (3) determining the supporting parameter scheme of the roof bolt of the main withdrawing channel by combining the technical requirements on the bolt supporting design of the GB35056-2018 coal mine roadway bolt supporting technical specification as follows: the phi 18 multiplied by 2000mm glass anchor rods have the row spacing of 800mm multiplied by 1000mm, and the mining side is not supported.
Establishing a numerical simulation:
respectively additionally arranging a real coal area and a goaf on two sides of an N1114 working face for simulation analysis, wherein the size of a model is X multiplied by Y multiplied by Z which is 300m multiplied by 650m multiplied by 150m, the length of a simulation stope working face is 200m, and one step is excavated every 20 m; when the tunnel is pushed to a distance of 50m from the retracting channel, excavating one step every 10 m; when the working face is 15m away from the mining stopping line, one step is excavated every 5m, and when the working face is 10m away from the main withdrawing channel, mining height reduction treatment is carried out, so that the unbalanced force is reduced to 5e-4The recovery rate is reduced in a simulated final recovery stage.
And (3) simulation result analysis:
wherein, the measurement of the change rule of the vertical stress is obtained by arranging a borehole stressometer in the withdrawal passage;
the main retrace channel roof subsidence data is obtained by arranging a convergence meter in the retrace channel.
FIG. 3 is a diagram of an original supporting scheme, wherein after a tunnel is excavated, a concrete surface layer is primarily sprayed and then is supported, a lead wire mesh 5 protects the surface, and a steel belt 4 is connected with a roof anchor cable 2; wherein the roof bolt 2 is a phi 20 multiplied by 2200mm left-handed non-longitudinal rib deformed steel bar bolt, and the row spacing is 800 multiplied by 800 mm; the top plate anchor cable 1 is made of a steel strand with the specification of 1 x 7-17.8-1860, namely the steel strand is made of a 1 x 7 structure steel strand, the nominal diameter is 17.8mm, the nominal tensile strength is 1860MPa, the length of the top plate anchor cable is 7000mm, the row spacing is 1600mm, and the spacing is 2350 mm; the upper part anchor rod 3 is a round steel anchor rod with phi 16 multiplied by 1800mm, the interval is 800mm multiplied by 800mm, the steel belt 4 is a phi 16 multiplied by 1800mmW type steel belt, the interval is 800mm multiplied by 800mm, and the lead wire mesh 5 is a 10# lead wire mesh; the simulation results of the original supporting scheme are shown in fig. 4 to 7.
The change rule of the original support vertical stress is as follows: from fig. 4, when the working face is advanced to about 28m from the withdrawal passage, the stress at the detection point of the mining side begins to be influenced by mining, the original rock stress is continuously increased from 3.17MPa, and the change rate shows an increasing trend. When the width of the residual coal pillar is reduced to 2m, the stress value of the mining slope measuring point reaches the maximum value of 5.97MPa, the stress concentration coefficient is 1.5, and then the residual coal pillar is subjected to plastic failure, so that the bearing capacity is reduced rapidly;
from fig. 5, when the working face is pushed to about 28m away from the withdrawal passage, the stress at the non-mining side measuring point begins to be influenced by mining, the original rock stress is continuously increased from 3.13MPa, and when the working face is pushed to 12.5m, the stress change rate tends to increase, which indicates that the advanced supporting pressure of the working face is transferred to the coal pillar at the moment. When the width of the residual coal pillar of the working face is 2.5m, the stress change rate at the measuring point of the non-mining side begins to be reduced until the working face is penetrated, the maximum value of the non-mining side is 6.71MPa, and the non-mining side is kept stable.
The sinking change rule of the top plate of the original support main withdrawing channel is as follows: from FIG. 6, when the working surface is pushed to about 50m away from the retracting channel, the vertical displacement at the measuring point is gradually increased; when the working face is about 12m away from the retraction channel, the sinking rate of the top plate is obviously increased; when the working face is communicated, the final sinking amount of the middle position of the retracting channel is 105 mm.
The convergence change rule of two sides of the original support main withdrawal channel is as follows: from fig. 7, when the working surface is pushed to about 36m away from the retraction channel, the horizontal displacement at the measuring point is gradually increased, and when the working surface is pushed to about 12m, the horizontal displacement change rate of the two sides is increased until the working surface is communicated and reaches the maximum value of 39.5 mm. The change of the horizontal displacement of the two sides in the change process has no obvious stable stage, because the displacement of one side of the mining side is influenced by mining and is changed violently until the working face is communicated, and the mining side is always in a horizontal moving state.
The simulation result of the support scheme determined by the method of the invention is analyzed: through comparison, the change rule of the vertical stress of the mining side and the non-mining side is basically the same as that of the original support in the working face advancing process.
The top plate sinking change rule of the supporting scheme determined by the method is as follows: from fig. 8, the maximum sinking amount of the top plate of the supporting scheme of the method determined by the invention is 53mm, which is reduced by 49.5% compared with the original supporting scheme, and the supporting scheme of the method determined by the invention can reduce the vertical deformation of the surrounding rock of the roadway, and can effectively control the deformation range of the surrounding rock, thereby weakening the influence of mining on two withdrawing channels.
The two-side convergence change rule of the supporting scheme determined by the method of the invention is as follows: from the figure 9, the convergence under the original supporting condition is 39.5mm, and in the supporting scheme of the method determined by the invention, the horizontal convergence of the two sides is 22.2mm, which is reduced by 17.3mm compared with the original supporting scheme, so that the horizontal displacement of the surrounding rock is effectively controlled, and the stability of the roadway is favorably maintained.
Step twelve, performing end mining and net hanging coal mining according to the sequence of coal cutting → net placing at 8-12 frames of a rear roller of the lagging coal mining machine → frame pulling → manual winch lifting and net lifting → pushing sliding → coal cutting;
thirteenth, when the working surface is 8.3m away from the withdrawal passage, a second steel wire rope is installed; when the working face is 6.3m away from the retraction channel, a third steel wire rope is installed; when the working face is 4.3m away from the retraction channel, a fourth steel wire rope is installed; when the working face is 3.3m away from the retraction channel, a fifth steel wire rope is installed; when the distance between the working surface and the retraction channel is 2.3m, a sixth steel wire rope is installed; when the distance between the working surface and the retraction channel is 1.3m, a seventh steel wire rope is installed;
fourteen, adjusting the coal mining rate according to the mine pressure observation data, and adjusting by combining the actual stoping condition of the working face; and determining to stop mining and withdraw the fully mechanized mining equipment according to the ore pressure analysis result.
The above description is only a preferred embodiment of the present invention, and is not intended to limit the present invention, and all simple modifications, changes and equivalent structural changes made to the above embodiment according to the technical spirit of the present invention still fall within the protection scope of the technical solution of the present invention.

Claims (10)

1. A fully mechanized coal mining face withdrawing and final mining method is characterized by comprising the following steps:
installing a bracket in the retraction channel;
adjusting the bottom plate of the working surface to be 180-220 mm higher than the bottom plate of the withdrawing channel;
determining support of an anchor rod and an anchor cable of the withdrawal channel, and then supporting the withdrawal channel by using the anchor rod and the anchor cable;
carrying out end mining and net hanging coal mining;
stopping mining and withdrawing the fully mechanized mining equipment.
2. The fully mechanized mining face withdrawal end mining method of claim 1, further comprising analyzing and predicting a periodic back pressure of a withdrawal path; the analyzing and predicting the periodic back pressure of the withdrawal passage comprises the following steps:
when the working face is recovered to 200m, carrying out first analysis and prediction on the periodic incoming pressure rule;
when the working face is recovered to 200-100 m, analyzing and predicting the periodic pressure-incoming rule every 50 m;
when the working face is recovered to be between 100m and 50m, the periodic pressure-incoming rule is analyzed and predicted every 25 m.
3. The fully mechanized mining face withdrawing end mining method of claim 2, wherein when the result of the first analysis prediction is through, the old roof is broken at an unfavorable position, and the mining height or the mining speed needs to be adjusted, and the result after the first measure is taken is evaluated in the second analysis prediction; the unfavorable position is right above the bracket; the support comprises a stack support or a hydraulic support.
4. The fully mechanized mining face withdrawing end mining method of claim 1, wherein the mounting stack support is between 150m and 100m from the face withdrawing to the main withdrawing channel.
5. The fully mechanized mining face withdrawing end mining method of claim 1, wherein the adjusting face floor is withdrawn at a face distance of 20m to 30m from the main withdrawing passage.
6. The fully mechanized mining face retraction end mining method of claim 1, wherein the method of determining support for the retraction channel bolt and cable includes the steps of:
step one, calculating the radius of a top plate fracture area and the radius of a top plate plastic area;
secondly, calculating parameters of the roof anchor rods, wherein the parameters of the roof anchor rods comprise the length of the roof anchor rods, the row spacing between the roof anchor rods and the prestress of the roof anchor rods;
calculating parameters of the top plate anchor cables, wherein the parameters of the top plate anchor cables comprise the length of the top plate anchor cables, the row distance between the top plate anchor cables and the prestress of the top plate anchor cables;
and step four, calculating parameters of the upper anchor rods, wherein the parameters of the upper anchor rods comprise the length of the upper anchor rods, the row spacing among the upper anchor rods and the prestress of the upper anchor rods.
7. The fully mechanized mining face withdrawing and final mining method according to claim 6, wherein the radius of the roof rupture zone and the radius of the roof plasticity zone in the first step are calculated according to the following formulas:
radius R of top plate crushing zonebThe calculation formula of (2) is as follows:
Figure RE-FDA0002385873090000021
the calculation formula of the radius of the plastic area of the top plate is as follows:
Figure RE-FDA0002385873090000022
wherein the content of the first and second substances,
Figure RE-FDA0002385873090000023
Figure RE-FDA0002385873090000024
Figure RE-FDA0002385873090000025
Figure RE-FDA0002385873090000026
Figure RE-FDA0002385873090000027
σcultimate compressive strength, MPa;
σb cthe compressive residual strength of the rock mass is MPa;
r0is the equivalent circle radius of the withdrawal passage, m;
q is the softening modulus of the strength of the roadway rock mass, GPa;
qithe strength of the collapse zone is MPa;
m is a correlation coefficient of the expansion fragmentation property of the surrounding rock plastic zone;
k is a plasticity coefficient;
P0is the stress of the original rock, MPa;
mu is Poisson's ratio;
σRis the radial stress of the elastic-plastic interface, MPa;
e is the modulus of elasticity, GPa.
8. The fully mechanized mining face withdrawing and final mining method according to claim 6, wherein in the second step, the parameter of the roof bolt is calculated according to the following formula:
roof anchor anchoring length LaThe calculation formula of (2) is as follows:
Figure RE-FDA0002385873090000031
wherein the content of the first and second substances,
k1the roof anchor rod anchoring force safety coefficient is set;
P1the roof anchor bearing capacity is kN;
d is the diameter of the anchor rod hole of the top plate, and is mm;
τrthe bonding strength between the anchoring agent and the rock mass is MPa;
the calculation formula of the roof bolt length L is as follows:
L＝La+a1+Lb； (4)
wherein the content of the first and second substances,
Lathe anchoring length of the roof bolt is m;
Lbthe exposed length of the roof bolt is m;
a1the depth m of the crushing area of the top plate of the withdrawing channel;
the calculation formula of the row spacing a between the roof anchor rods is as follows:
Figure RE-FDA0002385873090000032
wherein, γ1Is the coal bed severity, kN/m3；
G is the roof anchor rod anchoring force kN;
s1the safety factor of the row spacing between the roof anchor rods is set;
the calculation formula of the roof bolt prestress N is as follows:
Figure RE-FDA0002385873090000041
wherein gamma is the average top plate weight, kN/m3；
n is the distribution density of roof bolts, root/m2；
δ is the lower formation thickness, m;
α is the formation dip angle, °;
fais the interlayer friction coefficient.
9. The fully mechanized mining face withdrawing and final mining method according to claim 6, wherein the parameters of the roof anchor cable in step three are calculated according to the following formula:
anchoring length b of anchor cable on top plate2The calculation formula of (2) is as follows:
Figure RE-FDA0002385873090000042
wherein the content of the first and second substances,
K1the safety coefficient of the anchorage force of the anchor cable of the top plate is set;
P2the roof anchor cable bearing capacity, kN;
d is the diameter of the drilled anchor cable hole of the top plate, and is mm;
the calculation formula of the arch height b of the natural balance arch is as follows:
Figure RE-FDA0002385873090000043
wherein the content of the first and second substances,
b is the span of the withdrawal passage, m;
a2the upper plastic zone depth, m;
fbis the coefficient of Pyth; the length L of the roof anchor cabledThe calculation formula of (2) is as follows:
Ld＝b+b1+b2(9)
wherein the content of the first and second substances,
b is the natural balance arch height m;
b1the exposed length of the anchor cable of the top plate is m;
b2the anchoring length of the anchor cable of the top plate is m;
the calculation formula of the row spacing between the roof anchor cables is as follows:
Figure RE-FDA0002385873090000044
wherein the content of the first and second substances,
PNbreaking force is provided for each row of anchor cables; kN;
K3the safety factor of the row spacing between the anchor cables is set;
the roof anchor cable prestress N3The calculation formula of (2) is as follows:
Figure RE-FDA0002385873090000051
n3the anchor cables of the top plate are distributed with density of root/m2。
10. The fully mechanized mining face withdrawing and final mining method according to claim 6, wherein the parameter of the side anchor rod in the fourth step is calculated according to the following formula:
anchoring length L of anchor rod on side2The calculation formula of (2) is as follows:
Figure RE-FDA0002385873090000052
k2the safety coefficient of the anchoring force of the anchor rod at the upper part is set;
P4the bearing capacity of the side anchor rod is kN;
D4drilling holes for anchor rod holes on the upper part, wherein the diameter of the holes is mm;
the side anchor rod L1The length is calculated as:
L1＝L2+a4+L4(13)
wherein the content of the first and second substances,
L2the anchoring length of the anchor rod at the upper part is m;
L4the exposed length of the anchor rod at the upper part is m;
the calculation formula of the row spacing between the side anchor rods is as follows:
Figure RE-FDA0002385873090000053
wherein the content of the first and second substances,
s2the safety factor of the row spacing between the anchor rods at the upper part is set;
the calculation formula of the prestress of the upper anchor rod is as follows:
Figure RE-FDA0002385873090000054
wherein the content of the first and second substances,
n4the density of the anchor rod at the upper part is distributed, root/m2。
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