The failure mechanisms of engineering rock masses primarily involve tensile and shear failure. Differentiating between the acoustic emission (AE) signals generated during the tensile and shear damage processes in rock can provide a scientific basis for the classification of acoustic signals in field rock fracture monitoring. This paper presents a study on acoustic emission monitoring during the direct tensile testing of granite, proposing a method for classifying AE signals based on the damage and failure processes of the samples. Additionally, the classification of tensile and shear AE signals is explored. The main conclusions are as follows. The proportion of low-frequency signals (frequency <200 kHz) and high-frequency signals (frequency >200 kHz) in all AE signals was found to be 81.6 % and 19.4 %, respectively. Based on an integrated classification and statistical method for AE signals in rock tensile failure, which involves steps such as “denoising the raw waveform, time-frequency domain data transformation, fuzzification processing, extraction of dominant frequency and corresponding amplitude, and identification of secondary dominant frequencies,” the AE signals were categorized into two types, A and B. Type A signals accounted for an average of 7.6 %, while Type B signals made up 92.4 %. Based on the polarity determination method, the focal mechanisms of AE (Acoustic Emission) events were identified. In tensile events, the average proportion of Type A signals was 8.34 %, while the average proportion of Type B signals was 91.66 %. The Brazilian splitting test also yielded classification results similar to those obtained from direct tensile testing. Thus, it was preliminarily concluded that Type A signals, characterized by the presence of both a primary and secondary frequency, correspond to shear signals, whereas Type B signals, which only exhibit a primary frequency without a secondary frequency, correspond to tensile signals.

To gain an in-depth understanding of the failure characteristics and reinforcement mechanisms of the anchorage layer in tunnel, a series of loading failure experiments were conducted. By isolating the anchorage layer, its bearing capacity was effectively quantified. This study systematically examined the stress, deformation, and failure characteristics of the surrounding rock under unsupported conditions (surrounding rock layer, SRL), anchor bolt support (anchorage layer, AL), and pre-tensioned anchor bolt support (pre-tensioned anchorage layer, PAL). Furthermore, the incorporation of P-wave velocity (Vp) and PIV testing provided a robust framework for elucidating the reinforcement mechanisms of the anchorage layer. The failure of SRL initiated at the tunnel shoulders, whereas for AL and PAL, failure originated at the tunnel crown. The Vp at the tunnel crown of PAL and AL exhibited increases of approximately 11.0

The attainment of the pure shear state in rock presents a significant challenge for conventional laboratory testing methods. To achieve the pure shear state conveniently, a Double-Notch Shear method is developed in this study. Marble samples are tested using an MTS 815 rock mechanics test system and a PCI-2 acquisition system, and the feasibility was validated through the mechanical and acoustic emission (AE) results. To acquire the appropriate geometry of the sample, a double-notch mechanical method for numerical simulation is established. The main factors affecting the distribution of stress are identified; an investigation into the impact of notch width, notch depth, hollow cylinder thickness, circular table radius, and notch chamfering has been conducted using the ABAQUS numerical simulation method. Moreover, the optimal parameter thresholds have been obtained based on the refined stress analysis. The results show that the preferred depth falls within the range of 14–16 mm with an optimal width of 3–6 mm for the upper notch, while for the lower notch, depths of 14–16 mm and widths of 3–4 mm are considered optimal. The hollow cylinder thickness is considered suitable within the 40–49 mm range and the circular table radius is within the 35–44 mm range. In addition, implementing chamfering measures at the notched edge has been demonstrated to mitigate stress concentration phenomena effectively. The determined sample parameters serve as valuable references for laboratory testing protocols.

During the construction of long-distance high geothermal tunnels, the ducts will exchange heat with the air inside the tunnels, resulting in the temperature of the air flow inside the ducts will increase. However, many scholars ignore this factor, leading to small results in calculating the ambient temperature and cooling air supply. Therefore, a prediction method for ambient temperature and air supply volume of high geothermal tunnels is proposed based on three-dimensional difference and heat transfer principles. It takes into account the external temperature of the tunnel, the construction length, the air volume and the duct radius, and this study is compared with field tests and other algorithms to verify the accuracy. The conclusions of the study are as follows: The prediction method is compared with the measured ambient temperature, and the average ambient temperature difference is 0.1 °C, which verifies the accuracy of this study. The ambient temperature increases by 10.74 °C for every 1000 m increase in construction length. When the TISR is 60 °C and the TOUT is 10 °C, for every 1% increase in the air leakage rate, the air volume for cooling needs to increase by 5-51 m3/s. The results of the study are more in line with the field and provide a basis for long-distance construction ventilation and cooling of high geothermal tunnels.

Tunnels with surrounding rock temperatures exceeding 28°C are called high geothermal tunnels. During the construction of high geothermal tunnels, construction personnel health, machine operation efficiency and second lining safety are affected by the heat tunnel. Therefore, calculating the precise heat release characteristics of high-geothermal temperature tunnels is a prerequisite for ensuring the smooth passage of the tunnel. The guideline that states that the wall temperature is constant in the longitudinal direction of the tunnel is not realistic. This constant temperature results in large heat release calculations and leads to increased energy wastage while cooling high geothermal tunnels. Therefore, a method is proposed for calculating the heat release along the longitudinal direction of a tunnel based on a three-dimensional implicit finite-difference technique considering the ventilation time and average wind speed. This calculation method is used to determine the heat release values of the surrounding rock, initial support and second lining and to compare these findings with the guideline algorithm. The results show that the difference between the internal surrounding rock temperature calculated by the three-dimensional implicit finite-difference method and the measured value is 0.94 °C, verifying the theory accuracy. The heat release of the initial support section is much higher than that of the surrounding rock and the second lining; all heat release values are smaller than the normative heat release. Therefore, the heat release calculation method should be extended and used as a basis for calculating ventilation and cooling in other high geothermal tunnel construction projects.

Tunnels with surrounding rock temperatures exceeding 28 degrees C are called high geothermal tunnels. During the construction of high geothermal tunnels, construction personnel health, machine operation efficiency and second lining safety are affected by the heat tunnel. Therefore, calculating the precise heat release characteristics of high-geothermal temperature tunnels is a prerequisite for ensuring the smooth passage of the tunnel. The guideline that states that the wall temperature is constant in the longitudinal direction of the tunnel is not realistic. This constant temperature results in large heat release calculations and leads to increased energy wastage while cooling high geothermal tunnels. Therefore, a method is proposed for calculating the heat release along the longitudinal direction of a tunnel based on a three-dimensional implicit finite-difference technique considering the ventilation time and average wind speed. This calculation method is used to determine the heat release values of the surrounding rock, initial support and second lining and to compare these findings with the guideline algorithm. The results show that the difference between the internal surrounding rock temperature calculated by the three-dimensional implicit finite-difference method and the measured value is 0.94 degrees C, verifying the theory accuracy. The heat release of the initial support section is much higher than that of the surrounding rock and the second lining; all heat release values are smaller than the normative heat release. This method of calculating heat release is more realistic and can effectively lay the foundations for high geothermal tunnel cooling while saving costs. Therefore, the heat release calculation method should be extended and used as a basis for calculating ventilation and cooling in other high geothermal tunnel construction projects.

Injecting carbon dioxide CO2 into a coal seam is an important way to improve coalbed methane recovery and to store geological carbon. The fracture mechanical characteristics of bituminous coal determine the propagation and evolution of cracks, which directly affect CO2 storage in coal seams and the efficiency of resource recovery. This study applied CO2 adsorption and three-point bending fracture experiments using bituminous coal samples in a gaseous state (4 MPa), subcritical state (6 MPa), and supercritical state (8 and 12 MPa) to investigate the influence of CO2 state and anisotropy on the fracture-related mechanical response of bituminous coal. The results show that the change in mechanical properties caused by CO2 adsorption is CO2 state-dependent. The supercritical CO2 adsorption at 8 MPa causes the largest decrease in the mode-I fracture toughness (KIC), which is 63.6% lower than the toughness before CO2 adsorption. The instability characteristics of bituminous coal show the transformation trend of “sudden-gradual-sudden fracture”. With or without CO2 adsorption, the order of the KIC associated with three types of bituminous coal specimens is crack-divider type > crack-arrester type > crack-short transverse type. Phenomenologically, the fracture toughness of bituminous coal is positively correlated with its specific surface area and total pore volume; the toughness is negatively correlated with its average pore size.

Flowing sand is a special surrounding rock encountered by tunnel construction. Due to the looseness and low viscosity of the flowing sand, after excavation, the sand body is easy to flow along the open surface. In addition, the water seepage also causes tunnel instability. Considering the characteristics of water seepage, how to improve the stability of flowing sand bodies and prevent the instability of surrounding rocks has become a difficult problem. In this paper, a parametric experiment on the surrounding rock taken from the project site was carried out, and then, a numerical simulation of the flowing sand body was conducted to study the precipitation construction method and stability of the flowing sand body. Other than that, the tunnel face vacuum dewatering, vertical vacuum dewatering at the top of the tunnel, and the vacuum dewatering technology of the gravity well in poor geological section were systematically analyzed in our research. A radial vacuum enclosed precipitation process for the face of the tunnel was proposed, which effectively solved the problem concerning continuous seepage of water in the front. Through numerical simulation and field experiments, the basis for determining the precipitation parameters of the tunnel face was obtained, while aiming at the top position of the tunnel, a vertical vacuum negative pressure precipitation method of intercepting the top seepage water and the water supply behind the top of the tunnel was proposed. For the bottom of the tunnel, setting gravity wells on the side walls for the purpose of preventing seepage at the bottom was put forward. The application of these methods in the project ensured the safety of construction and improved the construction schedule. After the completion of the dewatering construction, the method of inserting plywood into the small pipe was adopted to avoid the collapse of the dry sand. Then, to solve the problem of borehole collapse in flowing sand bodies, pipe feeding was introduced, thus further enhancing the precipitation effect. Furthermore, in view of the problem that the dewatering hole in the flowing sand body is easy to collapse, resulting in the failure of 60% of the dewatering hole and the sand body is extracted from the dewatering pipe, causing the risk of the cavity at the top of the tunnel, a method of pipe following is presented to avoid the damage of geotextile caused by directly inserting the dewatering pipe and further improve the dewatering effect. All the above processes together form an omnidirectional three-dimensional negative pressure precipitation method that considers the special sand body flow and water seepage of unfavorable geology and that has been proved to enhance the stability of surrounding rock in practice.

Collapse of the vault and numerous other safety accidents often occur during the construction process of large-section tunnels. The utilization of a small pilot tunnel and a step reverse expansion construction methodology is proposed based on conventional construction methods to explore safe construction technology. First, a theoretical analysis combined with on-site monitoring parameters was conducted. It showed that the maximum displacement of the tunnel surrounding rock was 0.027 m during the elastic stage and increased to 0.031 m during the strength limit stage. The overall surrounding rock deformation does not have a noticeable impact on tunnel safety. A numerical simulation model of the small pilot tunnel advancement and step reverse expansion method was established. Simulation results showed that the first two excavation steps caused 89.6% of the total overlining strata subsidence, and the use of a small pilot tunnel advancement and step reverse expansion method can enhance the tunnel support. The tunnel surrounding rock was adequately stabilized after using this excavation method and provides the in-situ conditions for expanding the pilot tunnel to the large-section tunnel. The proposed method was adopted in an actual engineering project. It protected the subsequent construction of the main tunnel and decreased construction time, saving construction costs while ensuring safety, reducing construction risks, and improving production efficiency. This research can guide similar tunneling projects.

针对倾斜坚硬岩石地段地下连续墙及钻孔桩施工困难、容易偏孔等难题,文章以某地铁车站深基坑围护结构施工为研究对象,通过理论研究及施工实践提出了旋挖钻机成孔+冲孔桩机成槽施工技术以及潜孔钻机+旋挖钻机成孔施工技术.结果表明:1)在坚硬岩石地段施工地下连续墙时,采用旋挖钻机间隔施工主孔,采用冲孔桩机进行副孔施工,可提高地下连续墙施工速度;2)在坚硬岩石地段施工钻孔桩难度较大,采用潜孔钻在桩位先进行梅花形钻孔,再用旋挖钻进行施工,施工速度提高明显;3)通过对旋挖钻机及潜孔钻垂直度的调整,避免了成槽成孔的倾斜,保证了在倾斜岩体中成槽成桩的垂直度;4)在倾斜坚硬岩石地段采用旋挖钻施工时,为了保证成孔成槽的垂直度,需对旋挖钻机的技术参数进行优化.

研究目的:目前我国城市地铁正在迅猛发展,其中遇到的问题较多,本文以某框架结构七层办公楼为研究对象,针对现场施工左线隧道拱顶下沉及地面沉降速率超过预警的现象,理论分析结合MIDAS-GTS数值模拟的方法,将“建筑物-地层-隧道”作为研究对象,分析矿山法小净距隧道施工工序对临近建筑物影响,分析左右线施工顺序对其影响差异性,并提出相关工程技术措施.研究结论:(1)隧道施工使得建筑物处于受拉区,结构容易发生破坏,同时距离建筑物超出开挖影响范围时,小净距隧道开挖先后顺序影响不大;(2)距离建筑物较近,处于围岩破碎区时,先开挖远离建筑物一侧的隧道方案更优,同时要注意靠近建筑物一侧及中间岩柱需采取加固措施;(3)本研究结果可应用于地铁盾构施工领域,对盾构近距离侧穿建筑物具有一定的指导作用.

为验证特大断面硬岩隧道爆破空孔布置形式及爆破参数选取的合理性,并解决现场试爆造成的危险系数增加和费用增高等问题.对丰宁隧道Ⅱ级围岩段采用ANSYS/LS-DYNA有限元模拟软件,模拟了中间空孔和四周空孔2种布置形式的直眼掏槽方案,分析爆破应力波传播规律及岩石破坏情况,确定前者为优选方案;利用光面爆破技术,确定周边眼合理参数;结合现场观测数据,评价爆破效果.证明“中间空孔布置的直眼掏槽+合理的爆破参数”设计是合理的,且数值模拟技术可有效降低成本,其研究成果可为类似工程提供参考.

To study the roof anchor support for open-off cut in soft coal mechanized mining,the roofs which had been broken were processed as continuity in numerical simulation.A series of different anchor thicknesses were simulated.After the analysis of horizontal stress in the center of surface and vertical disposition field,the results were: when the thickness was small(width-thickness ratio 3/10),the interior tensile stress was the decisive factor in roof destruction;when the thickness was large(width-thickness ratio 3/10),the interior shear stress was.Therefore,by the theoretical analysis and numerical simulation,the rational ratio between the anchor thickness and open-off cut width is 3/10.

On the basis of the cross-section enlargement construction of coal tunnels in Shanzhuang mine,Xihe Coal Co.,Shanxi Coal Transportation and Sales Group,the FLAC3D numerical simulation was used to compare the surrounding rocks stress distribution in the different conditions as follows: with the original wooden-shed support,after the removal of wooden-shed support,after the cross-section enlargement,and with the different supporting sequence.The deformation damage features of roof and two sides in the construction were analyzed,and then the technique optimization plan was established,including roof excavation first and then two sides step-by-step excavation and step-by-step supporting used in excavation.

During the extraction process of the top coal caving working face, the fierce looseness, deformation as well as fracture of the surrounding rocks would usually occur because the slots will get strong disturbance. Meanwhile, anchored bolt and cable-bolt shall deform accordingly. When the deformation of the anchored bolt or cable-bolt comes to the extreme extent, they will crack. Then the catapulting and wounding incident will easily happen if the fractural part of the anchored bolt and cable-bolt is not fully griped by the rocks. In the paper, based on the dynamic mechanism analysis on fracture phenomenon about the anchored cable-bolt, the formulas of the kinetic energy and speed of the fractural catapult of the horizontally anchored cables in the ribs and the vertically anchored cables in the roof have been established. And a bolt support theory “small pretightening force & large expected anchorage strength” has been put forward, which is useful for seriously deformed roadways to prevent anchored cable-bolts from breaking. In addition, a technological method to stop anchored cable-bolt ejecting by means of pipe cap has been put forward.