新时代下要培养具备扎实专业知识的应用型人才和创新性人才,需要探索工程地质与水文地质课程更高效、实用、创新的教学方式.文章从工程地质的重要性、教学内容调整、教学方法方式探索、学期末考核方式调整等方面进行论述,以此提升学生的自主学习能力、科技创新能力以及团队协作能力,使工程地质专业培养既具有扎实的通用基础,又具有专业特色,使学生既具有从事不同行业工程地质的广泛适应能力,又兼顾特色培养下的创新能力和比较优势,提升学生专业素质和竞争能力.

倾斜煤层厚层坚硬顶板条件是东北地区煤矿开采面临的常见条件,以东保卫煤矿倾角25°、14 m厚砂岩顶板为研究对象,采用力学理论分析,推导出了倾斜煤层厚硬顶板切顶支护阻力参数计算公式;采用UDEC数值模拟发现,切顶高度增加可以在减小围岩变形量、减小煤帮峰值压力并促使应力峰值向煤壁深处转移等方面起到卸压作用;切缝角度增加,围岩变形呈先减小后增大、煤帮应力峰值大小以微弱变化向深部移动的规律;最后基于砌体块体稳定理论,得出切顶高度和角度的理论计算方法.将研究成果综合应用于工程中,确定了东保卫煤矿三采区-620左面工作面成巷切顶高度为7.4 m、切顶角度为20°、爆破参数装药结构为4+4+2、巷内支护阻力为2741 kN的参数设计值,以及分区支护范围数值,现场工业试验表明,留巷成型良好,可为同类条件切顶留巷设计及应用提供有益参考.

随着我国林业发展和对绿色建筑业的重视,现代木结构工程应用获得较快发展,而我国木结构相关的技术人员十分匮乏,亟须迅速提升高校木结构人才培养质量及教学质量.文章针对当前高校木结构教学、社会人才需求现状及存在问题,从学科交叉、工程案例剖析、双语教学、建筑竞赛结合等多途径进行教学实践探索,以提高木结构课程教学质量与综合性木结构人才培养水平.

In order to study the mechanical mechanism of the new tension–compression composite anchor, the finite element software is used to establish the numerical calculation model of the new composite anchor, and the numerical drawing test is carried out. Through numerical experiments, the optimal position range of the bolt bearing plate is obtained, and the axial load and shear stress distribution of the bolt were analyzed. The results show that the ultimate pullout force of the new tension–compression composite bolt increases first and then decreases with the movement of the bearing plate position, and the optimal position is at the position where the length ratio of the tension–compression anchorage section is 3:4. At the same time, the study shows that the optimal position of the bearing plate has an optimal position interval. The ultimate uplift capacity of the new tension–compression composite anchor is 118.6% and 94% higher than that of the traditional tension anchor and pressure anchor, respectively. The shear stress distribution of the new tension–compression composite anchor is more uniform. The peak shear stress first appears at the bearing plate and begins to transfer to both sides of the bearing plate with the increase in the drawing force, which increases first and then decreases. Since the new tension–compression composite anchorage section is in an unbonded state, the axial load of the bolt is always equal to the drawing force. The axial force distribution of the tensile anchorage section is similar to that of the traditional tensile bolt and decreases exponentially along the axial direction of the bolt. Under the action of ultimate drawing force, the plastic zone distribution of the new tension–compression composite anchor is different from that of the traditional tension anchor and the traditional pressure anchor. The plastic zone first appears in the position of the bearing plate and begins to transfer to both sides of the bearing plate with the increase in the drawing force. When the plastic strain exceeds the allowable strain of the interface, the bolt pulls out from the anchor hole and the bolt support structure fails.

Deformation control of soft rock in deep roadways has always been a challenge in coal mining. The strain softening and rheological properties of high-stress soft rock not only cause large deformation of the surrounding rocks, but also lead to a long duration of deformation. Based on previous research results obtained here in China as well as abroad, this study focused on the S2S2 mining roadway in the Xiaokang Coal Mine of Liaoning Province in China for research and studied the characteristics and damage of its surrounding rock deformation. Furthermore, the sectional form that should be used in roadway construction, as well as the support structure types and parameters were proposed to control the excessive deformation of the surrounding rock. It is suggested that a circular section should be adopted as the mining roadway section, and the high-strength rock should be bolted together, employing the combination of the anchor cable support method. Industrial tests have verified that the circular roadway under the new support mode can enable relative distance changes between the roadway walls, make the roof and floor relatively flat, and reduce the convergence speed to a value lesser than that of the original support mode. The most important advantage is that under this joint support mode, the deformation and damage of the circular roadway would be less affected by mining activity, so the deformation and damage of the mining roadway can be controlled effectively. Therefore, the research results of this study can provide a reference value for similar engineering projects.

Aiming at the large deformation instability problem caused by the excavation unloading of a coal roadway in deep-buried slowly inclined jointed rock mass, the geomechanical parameters and deformation failure characteristics of an engineering geomechanical model were investigated. The in-situ stress state of the model was measured with the stress relief method. The geological and mechanical properties of roadway surrounding rock were described. The surrounding rock structure was revealed with the electron microscopy scanning method, micro-fractures and randomly distributed joints highly developed in roadway surrounding rock. Field investigation and monitoring indicated the cross-section of roadway surrounding rock shrank continuously and the deformation distribution was obviously asymmetric. Shotcrete spalling and cable broken failures frequently occurred in the middle and ride side of roof and right rib. Based on the geomechanical conditions of the coal roadway, a discrete element numerical model of coal roadway in gently inclined jointed rock mass was established. The parameters of rock mass in the numerical model were calibrated. The model ran in unsupported condition to restore the evolution process of stress, crack propagation and deformation in roadway surrounding rock due to gradual deviatoric stress release caused by excavation. On this basis, the space-time evolution characteristics and law of stress, crack propagation and deformation were obtained and then the asymmetric large fragmentation and dilatation deformation failure mechanism of roadway surrounding rock in deep-buried slowly inclined jointed rock mass was revealed. The failure reasons of the support structure were analyzed, and the relevant support principles were proposed. The research results can provide scientific references for the stability control of roadways excavated in jointed rock mass.

In order to reduce large deformation failure occurrences in non-pillar longwall mining entries due to roof weighting behaviors, a case study in Halagou coal mine was conducted on optimization and control techniques for entry stability in non-pillar longwall mining. The Universal Discrete Element Code (UDEC) modeling was adopted to study entry stability in non-pillar mining, and the characteristics of deformation and stress and crack propagation were revealed. The large deformation transmission between the entry-immediate roof and the gob-immediate roof could be eliminated by optimizing the entry roof structure through a directional roof-cutting method. The localized tensile stresses generated in the entry-surrounding rock caused the generation of coalescent macroscopic fractures, which resulted in the instability of the entry. The tensile stress state could be inhibited by an active flexible support system through enhancing the confining pressure on the surrounding rock. Serious rotation subsidence occurs in the entry roof due to periodic weighting of the main roof, which could be greatly reduced by a passive rigid support pattern. The numerical and field test results both showed that the roof weighting pressure was offloaded by the technique and that the deformation of the entry surrounding the rock in non-pillar mining was quite small. Thus, the technique can effectively ensure the stability of the gob-side entry, which can provide references for entry stability control in non-pillar longwall mining.

在对白居寺进行现场工程地质勘查、地质雷达探测的基础上,得出了白居塔建筑形制和建筑结构特点,整个塔体位于高强度板岩为主的宗山山体上,承载力满足上部荷载需要.采用MIDAS GTS数值仿真软件,分析了自重荷载和地震作用下不同基岩高度的塔体的稳定性,模拟结果表明:(1)较高山体模型结构最大变位以6层墙体斜向下变位为主,较低山体模型整体结构变位以向中心的水平变位为主,自重作用下较高山体模型安全稳定系数高于较低山体,但建筑地基承载力均没问题.(2)在7度多遇地震作用下,两个模型整体结构的抗震稳定性较好,墙体最大剪切应变带出现在7层平台对应的内墙体下部;在8度多遇地震作用下,较高山体模型主体结构基本处于安全状态,较低山体模型主体结构不满足安全要求.

Fully mechanized roadway is an area of high accident incidence, and the problem is more prominent with the increase of mining depth of coal resources. The advance support scheme put forward for fully mechanized mining roadway with self-moving support and anchor combined unit, and the configuration and working principle of support robot was explicated. A determination method on the working resistance of advance support bracket was proposed based on mechanical coupling model of surrounding rock-advance support brackets. Taking the geological conditions of the Qishan mine as an example, the process of analyzing and determining the working resistance of the advance support bracket was described using this method. An advance support bracket was designed using SolidWorks, and the working load was obtained based on numerical calculations. Then the stress and strain were analyzed using ANSYS. The results showed that the designed advance support bracket can withstand the corresponding top plate pressure. It provides a new method and thought for study and development of advance supporting equipment in deep fully mechanized roadway.

To provide the mechanical data reference for the structural design and the control of the advance support brackets, the dynamic coupling model of the surrounding rock with advance support brackets was established by FLAC3D, the pressure variation law of surrounding rock and the influence on the support bracket was investigated in the deep coal roadway under two support conditions. The studied results show that under the joint support condition of advance support bracket with anchor bolt, the stress at the top of the bracket should be greater than 20 MPa, the maximum stress of 163 MPa at the vertical pillar, and the stress at the horizontal direction should be greater than 52.9 MPa. The results provide the basis for the design of advance support brackets for deep fully mechanized coal roadway.

Based on the characteristics that horizontal pressure between rocks is too small to form masonry beam after roof fracture in shallow-buried coal seam, the retaining gob-side entry fracture structure formed by falling down cutting roof.It was established that fracture mechanical model of lower slitting and upper bending cracks bilateral cantilever beam.The cutting-top working resistance was derived as the lateral basic roof fracture extends along the cracks of the cutting seam structure.The numerical model of surrounding rock structure under two fracture states of lateral basic roof was built.The movement and deformation law under two states of fracture structure indicated:When the cutting resistance is low, the roof breaks along the coal wall inside certain depth.The pour terraced cantilever beam of arc triangle direct roofs in emptied stope, can transfer large fracture dynamic load and overlying strata movement pressure along emptied roadway surrounding rock.With the unit cutting action of high shear resistance and cutting seam, the direct roof of caving zone and low basic roof were cut to slide along the cutting seam structure.It could reduce the structure length of transferring the overburden strata load from the side of arc triangle hanging deck, and optimize the roof bearing structure and surrounding rock stress, and also could reduce the additional load bore by the support body in the roadside by implementing active control of roof fracture state.At the same time, the hulking of waste rock could fill the empty area and make the overlying rock contact point move forward, and slow down the sinking and rotational deformation force of overlying rock.In the condition of the shallow seam and thin bedrock of Halagou mine, the observation of cutting large cross-section compounded roof experiment indicated that:advanced slitting on gob-side roadway can reduce support working resistance of lateral roof and immediate roof caving step distance.It could also reduce the roof fracture dynamic load caused by rock plate breakage.The actual cutting resistance dovetailed with the theoretical calculation results.

To reduce the coal pillar' s loss at thin and medium coal seam,a new technique,entry roof cutting,for gobside entry retaining without coal pillar has been put forward.That is to generate a structural plane of cutting seam by cumulative blasting between the roof of entry and the side roof of mined area in last working face,the structural plane may weaken the transfer of the mining face' s side roof movement to the surrounding rock and cut the roof along the cutting seam.The fallen waste rock separates the goaf and forms a gob,in addition,it fills the falling space to support the rock stratum above,finally,the gob-side entry retaining achieve its stability.The theory and numerical analysis of the mechanical model suggest that the cutting seam could change the rock beam structure,which could transfer the pressure of deformation of rock stratum,to the short cantilever rock beam structure which has low load and doesn' t transfer the pressure of deformation,thus could reduce the surrounding rock' s pressure and deformation of gob-side entry retaining.According to that principle,the technique has been put forward to combine reinforcing support in entry,supporting densely in the side of entry and the cutting seam of entry roof.The study shows a favorable effect in the project of roof cutting to form an entry within thin and medium coal seam in different conditions.The method could reduce the pressure from the move of rock stratum in mining and the deformation of surrounding rock effectively.

To deal with the problems associated with anchor rope support in tunnels with slow and instantaneous large deformation which occurred frequently in coal mines,an appropriate high constant resistance and large deformation anchor rope,which is made of the constant resistance device and anchor body,has been invented based on the oretical analysis,laboratory and field tests.The laboratory test showed that the constant resistance anchor rope provided a constant resistance of 350 kN and the displacement reached 300-950 mm under the static tension,and that the burst resisting force was 280-375 kN under the burst impact dynamic load when the limit elongation was not exceeding 1.6 m.the developed anchor rope displayed the detachment failure instead of breaking failure of traditional anchor rope.Once the anchor rope force exceeded the maximum constant value,the constant resistant cell began to slip.In the slipping stage,the temperature of the constant resistant cell increased rapidly first,then decreased slowly and approached stable finally.The highest working temperature is 35.6°.The external conjugate diameter of the constant resistant cellswasenlarged evenly along its axis,exhibiting a negative Poisson's ratio.The maximum value of radial expansion is 4.1 mm and the Poisson's ratio is between 0.5 and-2.0.In the blast impact failure test at backstopping tunnel,the high gas and constant resistance anchor rope entry retaining and the large section cut roof entry retaining,the maximum slipping value of constant resistant cell was 57 mm.The constant anchor rope has the advantage of low temperature,high pretightening force,and large elongation.

针对新维煤矿新场一盘区由于小煤矿存在,采空区众多,应力集中严重,巷道围岩为软岩,极为破碎,总回风巷已出现底臌、顶板下沉.本文运用松动圈研究理论,结合现场地质调查、超声波探测、钻孔窥视等方法,确定松动圈范围,并对巷道围岩情况进行评定,为总回风巷道返修及运输巷提供科学依据.

The roadway with soft rock has such problems as high ground pressure, strong rheology, and large deformation in deep coal mining in China, so the large supporting force is needed to stabilize the surrounding rock. The author has put forward a new technology with truss support, and the truss is made of eight 12# double I-Beams. Each beam connects another beam with 390 mm×800 mm×30 mm welded steel plate and bolted connection. In laboratory the truss mechanical property about ultimate load and deformation has been achieved by the monitoring system with different combination with horizontal load and vertical load that is stepwise added. The results have shown that under the uniform pressure ultimate bearing capacity of the truss is 1 500 t, and the ultimate deformation is 12.18 mm; when the ra-tio of the vertical load to the horizontal load is 0.2, ultimate load of the truss is 900 t, and the elastic de-formation limit is 26.3 mm; and the failure is the connection slide and the diameter shrinkage damage. Compared with the traditional U-shaped steel, the double truss support counterforce has been increased to 10 times. In the soft rock roadway, the author has put forward the first support with con-stant-resistance large-deformation anchor coupling support and reserved deformation; the support can release the deformation energy. After forming the anchor-surrounding rock coupling support circle, the strong truss force support has been used in the second support; this way can make full use of mechanical property. The truss structure may have an extensive application in fields of the deep roadway with the soft rock, high ground pressure and strong rheology in future.

采用杂交有限元-离散元法(FDEM),与日本等比尺浅埋隧道衬砌试验结果进行了对比分析.数值模型较好地再现了试验中衬砌结构性裂缝的初裂、增扩直至破坏的全过程.其次,建立荷载结构模型,针对影响衬砌开裂的3个主要荷载因素,即局部偏压、背后空洞和拱顶松弛地压,分别研究了这些压力不同作用位置、作用范围、作用大小下的裂缝扩展过程、分布规律.为用非连续方法研究衬砌裂缝开裂过程提供了借鉴,同时也为隧道衬砌裂缝病害的防治与加固修复打下了基础.

为在坚硬软弱复合顶板切顶卸压沿空留巷时取得较好预裂爆破效果,通过现场试验和分析,采用合理的双向聚能张拉爆破技术,对巷道顶煤强度较小、直接顶较软弱、基本顶坚硬的复合顶板进行了有效的预裂爆破.结果表明:对于软弱岩层,爆破能量易沿孔壁原生裂隙发生冲楔作用,聚能效果较差,深12m的炮孔内裂缝长度只有1.2m,并且孔口易出现爆破漏斗,对巷道原有支护造成破坏;对于坚硬岩层,随着装药线密度增大,聚能方向裂缝率逐渐增大,当单个炮孔装药量为7卷、药卷长度和空气柱长度的比值为3.348、填塞位置位于坚硬岩层时,普氏系数为12的中砂岩段炮孔内沿预裂方向全部产生张拉裂缝,裂缝长度达5 m,裂缝率达100％.因此,对于坚硬软弱复合复合顶板,使坚硬岩层沿预裂方向产生有效的裂缝,软弱岩层自行垮落,能够取得满意的预裂效果.

为解决小河嘴煤矿留煤垛护巷矿压显现剧烈的难题，文章对工作面回采后巷道顶板支承结构进行了分析，采用FLAC 2D数值软件模拟了普通爆破与定向聚能爆破效果，然后对小河嘴矿2012工作面风巷进行了巷旁留煤垛护巷与卸压短臂梁沿空留巷模拟对比研究，提出了小河嘴煤矿2018工作面运输巷卸压短臂梁沿空留巷工艺。

垮落带内含有厚层坚硬岩层时,难以冒落,易在沿空巷道采空区侧形成弧形三角悬板,对沿空巷道产生较大压力.以大同唐山沟8820厚层砂岩顶板首采面无煤柱开采为背景,分析普通充填留巷和切缝沿空成巷侧向顶板断裂结构特征及围岩稳定过程,认为对垮落带内直接顶坚硬层顶板进行合理参数下的切缝,可使得切缝高度范围内采空区边界直接顶和基本顶失去约束,并沿切缝结构面剪切破断充分冒落接顶,降低破断冲击动载;并通过UDEC数值模拟软件,分析出切缝有利于矸石冒落并支撑上覆岩层,可将上覆基本顶岩层的触矸点前移,限制基本顶回转和下沉作用引起的围岩压力,明显减小巷道围岩变形量.基于理论、数值分析研究结果,确定唐山沟矿8820回风巷巷内加强巷旁密集支柱+巷旁双向聚能爆破切缝的上压下支中间切缝的联合切顶方案.通过井下爆破参数试验、矿压监测分析,评价切顶成巷的效果.井下试验表明:顶板高恒阻大变形锚索、巷内加强巷旁密集支柱、巷旁密集档矸点柱、超前聚能切缝爆破的切顶成巷综合技术,能够有效切落沿空巷道侧向项板并形成完整巷道,各项指标均满足下一工作面回采要求.