Mechanisms and control of rock bursts in goaf-side roadways under the condition of thick and hard roofs in the Xinjie mining area, Inner Mongolia
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摘要:目的和方法
针对厚硬顶板条件下深部矿井回采巷道冲击地压严重威胁工作面安全生产的问题，以内蒙古新街矿区典型深采矿井3-1103工作面辅运巷为工程研究背景，分析邻空巷道冲击地压频发区域外在主控因素和内在驱动力源；构建基于软化地基与弹性地基假定“岩梁−地基”系统力学特性的顶板断裂前受载力学模型，解析采场覆岩结构演化过程厚硬顶板岩梁能量演化规律及其主控因素；运用FLAC3D模拟并探查邻空巷道冲击失稳高风险区域位置与特征；研究采场覆岩结构优化与围岩应力能量控制方案，制定厚硬顶板破断诱发邻空巷道冲击地压控制方法。
结果和结论结果表明：(1) 邻空巷道冲击地压频发区域易发生以高静载或高静载叠加动载为主导灾变力源的失稳破坏，影响因素主要为顶板厚硬岩层、邻近采空区、区段煤柱。(2) 顶板储能总量与覆岩载荷、软化地基系数、顶板岩梁弹性模量及惯性矩、采空区顶板极限跨距、工作面支架参数等有关。其中，覆岩载荷、软化地基系数和采空区顶板极限跨距与岩梁应变能密度呈正相关，顶板岩梁弹性模量及惯性矩、工作面支护参数与岩梁应变能密度呈负相关。(3) 回采期间3-1103工作面超前支承压力区及其影响区域内区段煤柱和回采巷道煤体呈现多因素叠加影响，发生应力集中和能量积聚，是冲击失稳高风险区域；该区域较3-1101综采面其围岩应力与能量集中度进一步加剧增大。其中，工作面前方应力和应变能密度峰值增幅最大分别为6.61%、12.04%，区段煤柱应力和应变能密度峰值增幅最大分别为29.06%、65.14%。(4) 提出了“卸压爆破预处理高静载区域+深孔爆破或水力致裂预裂厚硬顶板+强化巷道吸能防冲支护”的解决方案，现场应用效果明显。
Abstract:Objective and MethodsRock bursts in mining roadways in deep coal mines under the condition of thick and hard roofs severely threaten the safe coal mining of the mining face. To meet this threat, this study, focusing on the No.3-1103 mining face auxiliary transport roadway, a goaf-side roadway, in a typical deep mine of the Xinjie mining area, Inner Mongolia, analyzed the external dominant factors and internal driving sources for frequent rock bursts in the goaf-side roadway. Assuming the mechanical characteristics of the rock beam-foundation system, this study constructed mechanical models under loading before roof fracturing based on soft and elastic foundations. Using these models, this study determined the evolutionary patterns of the rock beam energy in the thick and hard roof during the structural evolution of overburden in the stope, as well as their dominant factors. Based on FLAC3D simulation results, this study investigated the locations and characteristics of high-risk zones of rock burst-induced instability in the goaf-side roadway. Finally, this study developed a scheme for optimizing overburden structures and controlling the stress energy of surrounding rocks in the stope, along with a method for controlling rock bursts in the goaf-side roadway induced by the fracturing of the thick and hard roof.
Results and ConclusionsThe results indicate that zones with frequent rock bursts in the goaf-side roadway are susceptible to instability failures caused primarily by high static loads or high static loads with superimposed dynamic loads. The primary influencing factors of such failures include the thick and hard roof, the adjacent goaf, and section coal pillars. The total energy storage capacity of the roof is related to factors like overburden load, soft foundation coefficient, the elastic modulus and moment of inertia of rock beams in the roof, the limit span of the goaf roof, and the support parameters of the mining face. Specifically, the strain energy density of the rock beams is positively correlated with the overburden load, soft foundation coefficient, and limit span of the goaf roof but negatively correlated with the elastic modulus and moment of inertia of rock beams in the roof and the support parameters of the mining face. During the coal mining along the No.3-1103 mining face, section coal pillars and coals in the mining roadway within the advance support pressure zone and its influencing zones are affected by superimposed multiple factors, which lead to stress concentration and energy accumulation. Therefore, these zones face high risks of rock burst-induced instability. Compared to the No.3-1101 fully mechanized mining face, these zones exhibit significantly intensified surrounding rock stress and energy concentration. The peak values of stress and strain energy densities in front of the mining face increase by 6.61% and 12.04% at most, respectively. In contrast, the peak values of stress and strain energy densities of section coal pillars increase by 29.09% and 65.14% at most, respectively. Finally, this study developed a comprehensive scheme involving the blasting and pressure relief of high static load zones, the pre-fracturing of the thick and hard roof through deep-hole blasting or hydraulic fracturing, reinforcing energy-absorbing and anti-rock burst supports in the roadway, with significant effects having been achieved in the field application of this scheme.
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图 1 工作面及巷道布置与3-1103辅运巷冲击显现区域
Fig. 1 Layout of mining faces and roadways and the zone with apparent rock bursts in the No.3-1103 auxiliary transport roadway
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图 3 坚硬顶板断裂前覆岩结构及受力分析模型
Fig. 3 Overburden structures before fracturing of the thick and hard roof and mechanical models under loading
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图 4 不同参数下顶板岩梁应变能密度分布规律 (初次来压)
注：m=6.69 m；Lk=5 m；Em=1.83 GPa；k1=Em/m。
Fig. 4 Distribution patterns of strain energy density of rock beams in the roof under different parameters during the initial weighting of mining face (first weighting)
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图 5 不同参数下顶板岩梁应变能密度分布规律 (周期来压)
注：m=6.69 m；Lk=5 m；Em=1.83 GPa；k1=Em/m。
Fig. 5 Distribution patterns of strain energy density of rock beams in the roof under different parameters during the periodic weighting of mining face (periodic weighting)
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图 6 3-1101综采面采场围岩应力三维空间分布
注：QⅠ区为3-1101工作面超前支承压力影响区域；QⅡ区为3-1101工作面采空区侧向支承压力影响区域内回采巷道；QⅢ区为邻空侧3-1101工作面端头区域；QⅣ区为3-1101工作面超前支承压力影响区域内回采巷道。
Fig. 6 The three-dimensional spatial distribution of surrounding rock stress in No.3-1101 mining face
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图 7 3-1103综采面采场围岩应力三维空间分布形态
注：QⅤ区为3-1103工作面超前支承压力影响区域；QⅥ区为3-1101与3-1103工作面间区段煤柱；QⅦ区为邻空侧3-1103工作面端头区域；QⅧ区为3-1103工作面超前支承压力影响区域内邻空侧回采巷道。
Fig. 7 Three-dimensional spatial distribution of surrounding rock stress in the No.3-1103 fully mechanized mining face
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图 8 不同长度顶板岩梁积聚应变能及其增幅变化曲线
Fig. 8 Curves showing the cumulative strain energy and its increased amplitude for rock beams under different roof lengths
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图 11 3-1101工作面回采期间卸压防冲方案
Fig. 11 Scheme for pressure relief and anti-rock burst during the coal mining along the No.3-1101 mining face
表 1 不同长度顶板岩梁应变能值与增幅变化
Table 1 Strain energy values and their increased amplitude for rock beams under different roof lengths
初次来压 周期来压 顶板长度L/m 顶板岩梁应变能Uc/MJ 顶板岩梁应变能增幅kc/% 顶板长度L/m 顶板岩梁应变能Uz/MJ 顶板岩梁应变能增幅kz/% 15 28.064 10 22.452 20 65.670 134.00 13 47.036 109.49 25 138.485 393.46 16 85.823 282.25 30 267.031 851.50 19 145.940 550.01 
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