CN115081073A - Dynamic pressure bearing support design method for non-pillar self-entry mining - Google Patents
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Abstract
The invention discloses a dynamic pressure bearing support design method for pillar-free self-entry mining, which relates to the technical field of mining and comprises the following steps: s1, determining a dynamic pressure bearing and supporting action mechanism of the coal pillar-free self-entry by combining the motion process and the breaking characteristics of the top plate of the short-arm beam, and further determining a temporary supporting range according to the dynamic pressure bearing and supporting action mechanism; s2, establishing a fracture mechanics model of the short-arm beam top plate under the action of dynamic pressure bearing support and considering influence factors, wherein the influence factors at least comprise top plate lithology and asymmetric support; s3, according to the nonlinear strength failure criterion and the upper limit analysis theory of the rock mass, combining the fracture mechanics model of the short-arm beam top plate, and providing a dynamic pressure bearing support design method. The method can explore the influence rule of different factors on the dynamic pressure bearing supporting force, and has important significance on the non-pillar self-entry mining theory and technical system.
Description
Technical Field
The invention relates to the technical field of mining industry, in particular to a dynamic pressure bearing support design method for pillar-free self-entry mining.
Background
The roof cutting pressure relief coal pillar-free self-entry technology is a new coal mining method without digging an entry in advance and reserving coal pillars, cuts off the stress transfer between a roadway top plate and a goaf top plate through a top plate directional joint cutting technology, effectively improves the stress environment of roadway surrounding rocks, and simultaneously fills the goaf by utilizing the mine pressure work and the rock mass crushing and swelling characteristic to automatically form a stoping roadway. The coal pillar-free self-entry technology can greatly improve the resource recovery rate, eliminate the harm caused by surrounding rock stress concentration due to the reserved coal pillars and has remarkable positive effect on the sustainable development of coal mining.
When no coal pillar is mined in the self-entry way, the self-entry roof and the broken stone slope in the area which is not stabilized yet behind the working face are continuously influenced by overlying strata movement and gangue collapse disturbance of the goaf, and the stage is a dynamic pressure bearing area and is a key supporting section of the entry. Numerous scholars research and apply dynamic pressure bearing support technology and form under different geological conditions. The existing research and application greatly promote the development of coal pillar-free mining theory and technology, but the design of a temporary supporting scheme for a dynamic pressure bearing area lacks a specific theoretical basis, and the supporting form and parameters are usually determined by virtue of field experience.
Therefore, a novel dynamic pressure bearing support design method for pillar-free self-entry mining is provided to solve the problems in the prior art.
Disclosure of Invention
The invention aims to provide a design method for dynamic pressure bearing support in coal pillar-free self-entry mining, which aims to solve the problems in the prior art, can explore the influence rule of different factors on the dynamic pressure bearing support force and has important significance on coal pillar-free self-entry mining theory and technical system.
In order to achieve the purpose, the invention provides the following scheme:
the invention provides a dynamic pressure bearing support design method for pillar-free self-entry mining, which comprises the following steps:
s1, determining a coal-pillar-free self-entry dynamic pressure bearing support action mechanism by combining the motion process and the breakage characteristics of the short-arm beam top plate, and further determining a temporary support range according to the coal-pillar-free self-entry dynamic pressure bearing support action mechanism;
s2, establishing a fracture mechanics model of the short-arm beam top plate under the dynamic pressure bearing supporting effect and considering influence factors, wherein the influence factors at least comprise overburden vertical stress, mining stress concentration coefficient, short-arm beam top plate thickness, roadway width, in-roadway anchor cable support, crushed gangue support coefficient, short-arm beam top plate lithology and asymmetric support;
and S3, calculating the dynamic pressure bearing support force according to the nonlinear strength failure criterion and the upper limit analysis theory of the rock mass and the fracture mechanics model of the short arm beam top plate.
Preferably, in the step S1, the coal-pillar-free self-entry dynamic pressure bearing support action mechanism includes controlling the fracture position of the short-arm beam top plate, and if the dynamic pressure bearing support strength is insufficient or the support is not timely, the short-arm beam top plate fractures along the position above the roadway; when the dynamic pressure bearing support strength in the roadway is enough, the roof-cutting short arm beam is broken above the lateral coal body, and the roof-cutting short arm beam is a target state to be achieved by the design of the dynamic pressure bearing support strength.
Preferably, in step S1, the determining the temporary support range according to the mechanical model and the theoretical calculation analysis includes the following steps:
s101: according to the rock sample obtained on site, measuring the residual crushing expansion coefficient K after the rock pieces are stably collapsed 1 ；
S102: according to the formula h s ＝h+h 1 -K 1 h determines keyRotary sinking h of block B s (ii) a Wherein h is 1 For coal mining height, h is the thickness of the short-arm beam at the top cutting, h s Is the rotary subsidence of the key block B, K 1 The coefficient of residual crushing and swelling of the waste rock is measured in a laboratory;
s103: according to the formula
Determining the length of a key block body B; wherein L is the length of the key block B, K 2 Is the fracture coefficient of the rock formation (related to the thickness, lithology and production rate of the key layer, and can be obtained by numerical simulation and model test), L s Is the periodic pressure step distance, h k Is the thickness of the critical formation, R T In order to control the tensile strength of the rock mass of the key layer, q is the overlying load borne by the rock beam of the key layer;
s104: taking the rock mass in the rotary deformation damage range as a rigid body, and taking alpha to be approximately equal to alpha' according to a formula
Determining deformation h of side roadway roof of mining area under stable state of rock caving s ' when the actually monitored deformation reaches h s Removing the temporary support to be used as a criterion of the time for withdrawing the temporary support;
wherein alpha is the rotation angle of the key block B of the basic roof, alpha' is the rotation angle of the deformation of the roadway roof, h s ' is the deformation of the side roadway roof of the goaf under the stable state of rock caving, and l is the roadway width.
Preferably, in step S2, when the fracture mechanics model of the short-arm beam top plate is established, a mining induced stress concentration coefficient and a support coefficient of crushed gangue are determined.
Preferably, the mining stress concentration coefficient is lambda 1 And λ 1 ＞1；
When the goaf rock mass does not generate reliable supporting force on the goaf roof, the load q borne by the key block is completely borne by the short-arm beam structure, and the bearing length of the upper part of the short-arm beam structure is L b ，L b ＝l+htanθ，
Wherein θ is the kerf angle.
Preferably, the support coefficient of the crushed expanded gangue is lambda 2 And 0 < lambda 2 ＜1；
Wherein λ is 2 ＝λ 1 K 3 cosθ，K 3 The side pressure conversion coefficient of the crushed expansive waste rock is influenced by the physical mechanical properties of the rock mass, the shape, the gradation and the arrangement mode of the collapsed rock mass, and can be obtained through numerical simulation and physical model compression tests of the crushed expansive rock mass.
Preferably, in step S3, according to the nonlinear strength failure criterion and the upper limit analysis theory of the rock mass, in combination with the fracture mechanics model of the short-arm beam top plate, the internal energy dissipation rate of the surrounding rock is solved, the external force acting power of the fracture mechanics model is determined, the dynamic pressure bearing support force is calculated, and a dynamic pressure bearing support design method is provided.
Preferably, the dynamic pressure bearing support force is calculated according to dynamic pressure bearing support parameters; the dynamic pressure bearing support parameters comprise overburden vertical stress, mining induced stress concentration coefficient, short-arm beam top plate thickness, roadway width, rock mass volume weight, rock mass cohesion, rock mass internal friction angle, broken swelling waste rock support coefficient and asymmetric coefficient.
Preferably, the formula for calculating the dynamic pressure bearing support force is as follows:
wherein F is dynamic pressure bearing supporting force, Pa is pre-tightening force required by a single anchor cable, l is roadway width, c is rock mass cohesion, h is roof-cutting short-arm beam thickness, and gamma is rock mass weight,
the internal friction angle of the rock mass, theta is a joint cutting angle, q is the overlying strata vertical stress, n is the number of anchor cables, k is an asymmetric coefficient, and lambda is 1 For exploiting the stress concentration coefficient, λ 2 Supporting coefficient of crushed and expanded gangue.
Preferably, according to the upper limit analysis theory, the determination criterion is to determine the dynamic pressure bearing support force under the critical condition when the roadway width l is equal to the caving body width b.
Preferably, the dynamic pressure bearing support force is asymmetrically arranged in the roadway.
Preferably, the asymmetric coefficient of the dynamic pressure bearing supporting force is k, and k is 1-1.4 obtained according to the stress and deformation monitoring of the supporting structure in the actual engineering.
Compared with the prior art, the invention has the following technical effects:
the invention provides a design method of dynamic pressure bearing support parameters by establishing a roof fracture mechanical model of a roof of a short-arm beam with a top cut and applying a rock nonlinear strength failure criterion and an upper limit analysis theory based on the motion process and the fracture characteristics of a roof of a coal-pillar-free self-entry. The research result has important significance for supporting equipment model selection, entry retaining parameter design, entry retaining process optimization and perfection of coal pillar-free self-entry mining theory and technical system.
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In order to more clearly illustrate the embodiments of the present invention or the technical solutions in the prior art, the drawings needed in the embodiments will be briefly described below, and it is obvious that the drawings in the following description are only some embodiments of the present invention, and it is obvious for those skilled in the art to obtain other drawings without creative efforts.
FIG. 1 is a mechanical model diagram of a pillar-free self-roadway mining roof structure in an embodiment of the invention;
description of reference numerals: A. b, C is a substantially top block.
Detailed Description
The technical solutions in the embodiments of the present invention will be described clearly and completely below, and it is obvious that the described embodiments are only a part of the embodiments of the present invention, and not all of the embodiments. All other embodiments, which can be derived by a person skilled in the art from the embodiments given herein without making any creative effort, shall fall within the protection scope of the present invention.
The invention aims to provide a design method for dynamic pressure bearing support in coal pillar-free self-entry mining, which aims to solve the problems in the prior art, can explore the influence rule of different factors on the dynamic pressure bearing support force and has important significance on coal pillar-free self-entry mining theory and technical system.
In order to make the aforementioned objects, features and advantages of the present invention comprehensible, embodiments accompanied with figures are described in further detail below.
As shown in fig. 1, the present embodiment provides a design method for dynamic pressure bearing support in pillar-free self-entry mining, which includes the following steps:
s1, determining a coal-pillar-free self-roadway dynamic pressure bearing and supporting action mechanism according to a test (a physical model test of an exploitation process of an N00 construction method) and field monitoring by combining a motion process and a breakage characteristic of a short-arm beam top plate, and further determining a temporary supporting range according to the coal-pillar-free self-roadway dynamic pressure bearing and supporting action mechanism;
s2, establishing a fracture mechanics model of the short-arm beam top plate under the action of dynamic pressure bearing support and considering influence factors, wherein the influence factors at least comprise overburden vertical stress, mining stress concentration coefficient, short-arm beam top plate thickness, roadway width, in-roadway anchor cable support, broken expansion gangue support coefficient, short-arm beam top plate lithology and asymmetric support;
s3, calculating dynamic pressure bearing and supporting force according to the nonlinear strength failure criterion and upper limit analysis theory of rock mass ((I) aiming at the nonlinear failure property of rock mass, Hoek provides a famous Hoek-Brown criterion internationally, which is a well-known basic theory in the field; and II) upper limit analysis theory, in plastic mechanics, assuming that the displacement state of a plastic deformation area is a dynamic-tolerant speed field, because the set speed field only needs to meet the dynamic-tolerant condition but not to consider the condition in stress, the speed field is not necessarily a real speed field.
In this embodiment, in step S1, the mechanism of the dynamic pressure bearing support action for coal-pillar-free self-entry includes controlling the fracture position of the short-arm beam top plate, and if the dynamic pressure bearing support strength is insufficient or the support is not timely, the short-arm beam top plate fractures along the position above the entry; when the dynamic pressure bearing support strength in the roadway is enough, the roof-cutting short arm beam is broken above the lateral coal body, and the roof-cutting short arm beam is a target state to be achieved by the design of the dynamic pressure bearing support strength.
In this embodiment, in the step S1, determining the temporary support range according to the mechanical model and the theoretical calculation analysis includes the following steps:
s101: according to the rock sample obtained on site, measuring the residual crushing expansion coefficient K after the rock pieces are stably collapsed 1 ；
S102: according to the formula h s ＝h+h 1 -K 1 h determining the rotary sinking h of the key block B s (ii) a Wherein h is 1 For coal mining height, h is the thickness of the short-arm beam at the top cutting, h s Is the rotary subsidence of the key block B, K 1 The coefficient of residual crushing and swelling of the waste rock is measured in a laboratory;
s103: according to the formula
Determining the length of a key block body B; wherein L is the length of the key block B, K 2 Is the fracture coefficient of the rock formation (related to the thickness, lithology and production rate of the key layer, and can be obtained by numerical simulation and model test), L s Is the periodic pressure step distance, h k Is the thickness of the critical formation, R T In order to control the tensile strength of the rock mass of the key layer, q is the overlying load borne by the rock beam of the key layer;
s104: taking the rock mass in the rotary deformation damage range as a rigid body, and taking alpha to be approximately equal to alpha' according to a formula
Determining deformation h of side roadway roof of mining area under stable state of rock caving s ' when the actually monitored deformation reaches h s Removing the temporary support to be used as a criterion of the time for withdrawing the temporary support;
wherein alpha is the rotation angle of the key block B of the basic roof, alpha' is the rotation angle of the deformation of the roadway roof, h s ' is the deformation of the side roadway roof of the goaf under the stable state of rock caving, and l is the roadway width.
In this embodiment, in step S2, when the fracture mechanics model of the short-arm beam top plate is established, the mining-induced stress concentration coefficient is determined in consideration of the mining influence of the working face, and the support coefficient of the crushed expansive gangue is determined in consideration of the support effect of the goaf gangue on the short-arm beam.
In this embodiment, the mining stress concentration coefficient is λ 1 And λ 1 ＞1；
When the goaf rock mass does not generate reliable supporting force on the goaf roof, the load q borne by the key block is completely borne by the short-arm beam structure, and the bearing length of the upper part of the short-arm beam structure is L b ，L b ＝l+htanθ，
Wherein θ is the kerf angle.
In this embodiment, the support coefficient of the crushed expanded gangue is λ 2 And 0 < lambda 2 ＜1；
Wherein λ is 2 ＝λ 1 K 3 cosθ，K 3 The side pressure conversion coefficient of the crushed expansive waste rock is (influenced by physical mechanical properties of the rock mass, shape, gradation and arrangement mode of the collapsed rock mass, and can be obtained by numerical simulation and physical model compression test of the crushed expansive rock mass).
In this embodiment, in step S3, according to the nonlinear strength failure criterion and the upper limit analysis theory of the rock mass, the fracture mechanics model of the short-arm beam top plate is combined, the internal energy dissipation rate of the surrounding rock is solved, the external force acting power of the fracture mechanics model is determined, the dynamic pressure bearing support force is calculated, and a dynamic pressure bearing support design method is provided.
Wherein, solving the internal energy dissipation rate of the surrounding rock comprises:
according to the nonlinear Hoek-Brown strength criterion of the rock mass and the associated flow rule thereof, the plastic positive strain rate and the plastic shear strain rate on the fracture surface of the rock mass and the normal stress and the tangential stress on the thin deformation layer of the fracture surface can be obtained;
and obtaining the internal energy dissipation rate of the unit volume at the fracture surface of the surrounding rock according to the parameters.
Determining the model external force acting power comprises the following steps: determining 6 parts of dead weight acting power of a short-arm beam top plate rock body, tunnel surrounding rock stress acting power, dynamic pressure bearing supporting force acting power, top plate anchor cable pretightening force acting power, lateral coal body supporting force acting power, crushed expanded gangue supporting force acting power and the like.
In this embodiment, the dynamic pressure bearing support force is calculated according to dynamic pressure bearing support parameters; the dynamic pressure bearing support parameters comprise overburden vertical stress, mining induced stress concentration coefficient, short-arm beam top plate thickness, roadway width, rock mass volume weight, rock mass cohesion, rock mass internal friction angle, broken swelling gangue support coefficient, asymmetric coefficient and the like.
In this embodiment, the formula for calculating the dynamic pressure bearing support force is as follows:
wherein F is dynamic pressure bearing supporting force, Pa is pre-tightening force required by a single anchor cable, l is roadway width, c is rock mass cohesion, h is roof-cutting short-arm beam thickness, and gamma is rock mass weight,
the internal friction angle of the rock mass, theta is a joint cutting angle, q is the overlying strata vertical stress, n is the number of anchor cables, k is an asymmetric coefficient, and lambda is 1 For exploiting the stress concentration coefficient, λ 2 Supporting coefficient of crushed and expanded gangue.
In this embodiment, according to the upper limit analysis theory, when it is determined that the width l of the roadway is equal to the width b of the caving body, the dynamic pressure bearing support force under the critical condition is determined.
In this embodiment, the dynamic pressure bearing support force is asymmetrically arranged in the roadway, wherein the asymmetry coefficient of the dynamic pressure bearing support force is k.
In this embodiment, the asymmetry factor is 1-1.4.
The principle and the implementation mode of the present invention are explained by applying specific examples in the present specification, and the above descriptions of the examples are only used to help understanding the method and the core idea of the present invention; meanwhile, for a person skilled in the art, according to the idea of the present invention, the specific embodiments and the application range may be changed. In view of the above, the present disclosure should not be construed as limiting the invention.
Claims (10)
1. A dynamic pressure bearing support design method for coal pillar-free self-entry mining is characterized by comprising the following steps:
s1, determining a coal-pillar-free self-entry dynamic pressure bearing support action mechanism by combining the motion process and the breakage characteristics of the short-arm beam top plate, and further determining a temporary support range according to the coal-pillar-free self-entry dynamic pressure bearing support action mechanism;
s2, establishing a fracture mechanics model of the short-arm beam top plate under the action of dynamic pressure bearing support and considering influence factors, wherein the influence factors at least comprise overburden vertical stress, mining stress concentration coefficient, short-arm beam top plate thickness, roadway width, in-roadway anchor cable support, broken expansion gangue support coefficient, short-arm beam top plate lithology and asymmetric support;
and S3, calculating the dynamic pressure bearing support force according to the nonlinear strength failure criterion and the upper limit analysis theory of the rock mass and the fracture mechanics model of the short arm beam top plate.
2. The design method for dynamic pressure bearing support in coal-pillar-free auto-entry mining according to claim 1, wherein in step S1, the mechanism of action of the coal-pillar-free auto-entry dynamic pressure bearing support includes controlling the fracture position of the short-arm beam roof, and if the dynamic pressure bearing support is not strong enough or is not supported in time, the short-arm beam roof is fractured along the position above the roadway; when the dynamic pressure bearing support strength in the roadway is enough, the roof-cutting short arm beam is broken above the lateral coal body, and the roof-cutting short arm beam is in a target state to be achieved by the design of the dynamic pressure bearing support strength.
3. The method for designing a dynamic pressure bearing support for pillar-free self-entry mining according to claim 1, wherein the step S1 of determining the temporary support range includes the following steps:
s101: according to the rock sample obtained on site, measuring the residual crushing expansion coefficient K after the rock pieces are stably collapsed 1 ；
S102: according to the formula h s ＝h+h 1 -K 1 h determining the rotary sinking h of the key block B s (ii) a Wherein h is 1 For coal mining height, h is the thickness of the short-arm beam at the top cutting, h s Is the rotary subsidence of the key block B, K 1 The coefficient of residual crushing and swelling of the waste rock is measured in a laboratory;
s103: according to the formula
Determining the length of a key block body B; wherein L is the length of the key block B, K 2 Is the fracture coefficient of the rock formation, L s Is the periodic pressure step distance, h k Is the thickness of the critical formation, R T In order to control the tensile strength of the rock mass of the key layer, q is the overlying load borne by the rock beam of the key layer;
s104: taking the rock mass in the rotary deformation damage range as a rigid body, and taking alpha to be approximately equal to alpha' according to a formula
Determining deformation h of side roadway roof of mining area under stable state of rock caving s ' when the actually monitored deformation reaches h s Removing the temporary support to be used as a criterion of the time for withdrawing the temporary support;
whereinAlpha is the rotation angle of the key block B of the basic roof, alpha' is the rotation angle of the deformation of the roadway roof, h s ' is the deformation of the side roadway roof of the goaf under the stable state of rock caving, and l is the roadway width.
4. The design method for the dynamic pressure bearing support for the pillar-free self-entry mining of claim 3, wherein in the step S2, when a fracture mechanics model of the short-arm beam top plate is established, a mining stress concentration coefficient and a support coefficient of crushed expanded gangue are determined.
5. The design method of the dynamic pressure bearing support for pillar-free self-entry mining of claim 4, wherein the mining stress concentration coefficient is λ 1 And λ 1 ＞1；
When the goaf rock mass does not generate reliable supporting force on the goaf roof, the load q borne by the key block is completely borne by the short-arm beam structure, and the bearing length of the upper part of the short-arm beam structure is L b ，L b ＝l+h tanθ，
Wherein θ is the kerf angle.
6. The design method for dynamic pressure bearing support in coal-pillar-free self-entry mining according to claim 5, wherein the support coefficient of the crushed expansive waste rock is λ 2 And 0 < lambda 2 ＜1；
Wherein λ is 2 ＝λ 1 K 3 cosθ，K 3 The lateral pressure conversion coefficient of the crushed and expanded gangue.
7. The design method for the dynamic pressure bearing support for the coal-pillar-free self-entry mining according to claim 2 or 3, wherein in the step S3, according to the nonlinear strength failure criterion and the upper limit analysis theory of the rock mass, the internal energy dissipation rate of the surrounding rock is solved by combining the fracture mechanics model of the short-arm beam top plate, the external force of the fracture mechanics model is determined to do work, the dynamic pressure bearing support force is calculated, and the design method for the dynamic pressure bearing support is provided.
8. The design method of dynamic pressure bearing support for coal pillar-free auto-entry mining according to claim 5, wherein the dynamic pressure bearing support force is calculated according to dynamic pressure bearing support parameters; the dynamic pressure bearing support parameters comprise overburden vertical stress, mining induced stress concentration coefficient, short-arm beam top plate thickness, roadway width, rock mass volume weight, rock mass cohesion, rock mass internal friction angle, broken swelling gangue support coefficient and asymmetry coefficient;
the formula for calculating the dynamic pressure bearing and supporting force is as follows:
wherein F is dynamic pressure bearing supporting force, Pa is pre-tightening force required by a single anchor cable, l is roadway width, c is rock mass cohesion, h is roof-cutting short-arm beam thickness, and gamma is rock mass weight,
the internal friction angle of the rock mass, theta is a joint cutting angle, q is the overlying strata vertical stress, n is the number of anchor cables, k is an asymmetric coefficient, and lambda is 1 For exploiting the stress concentration coefficient, λ 2 Supporting coefficient of crushed and expanded gangue.
9. The design method of the dynamic pressure bearing support for the coal pillar-free self-entry mining according to claim 1, wherein according to the upper limit analysis theory, the determination criterion is to determine the dynamic pressure bearing support force under the critical condition when the width l of the roadway is equal to the width b of the caving body.
10. The design method for dynamic pressure bearing support in coal-pillar-free self-entry mining according to claim 8, wherein the dynamic pressure bearing support force is asymmetrically arranged for a short-arm beam top plate in an entry;
and the asymmetric coefficient of the dynamic pressure bearing support force is k, which is obtained by monitoring the stress and deformation of the support structure in actual engineering, and k is 1-1.4.
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| CN116756838A (en) * | 2023-08-21 | 2023-09-15 | 北京科技大学 | Control method for self-forming roadway roof structure without coal pillar |
| CN117948143A (en) * | 2024-03-26 | 2024-04-30 | 煤炭科学技术研究院有限公司 | Roof fracturing method for coal mine working face in initial mining stage |
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| Publication number | Priority date | Publication date | Assignee | Title |
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| CN116756838A (en) * | 2023-08-21 | 2023-09-15 | 北京科技大学 | Control method for self-forming roadway roof structure without coal pillar |
| CN116756838B (en) * | 2023-08-21 | 2023-11-10 | 北京科技大学 | Control method for self-forming roadway roof structure without coal pillar |
| CN117948143A (en) * | 2024-03-26 | 2024-04-30 | 煤炭科学技术研究院有限公司 | Roof fracturing method for coal mine working face in initial mining stage |
| CN117948143B (en) * | 2024-03-26 | 2024-05-31 | 煤炭科学技术研究院有限公司 | Roof fracturing method for coal mine working face in initial mining stage |
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