The excavation unloading effect results in the unloading damage of rock，and the mechanical properties of unloading damage rock under load directly influence the stability and safety of engineering rock. Based on this，the unloading damage test of limestone and the reloading test of damaged rock samples are designed and carried out. The results show that：(1) The initial unloading damage has a significant impact on the deformation and bearing capacity of rock sample. When the unloading amount is greater than 80%，the compressive strength，deformation modulus and strain confining compliance show a nonlinear mutation trend. (2) On the basis of the characteristics of rock transformation from stable failure to unstable failure in the process of loading and unloading，the proportion of dissipated energy is proposed to determine the critical damage state of rock sample during unloading，the proportion of dissipated energy is about 0.33 as the limit. When the proportion of dissipated energy is less than 0.33，the initial unloading damage has little influence on its reloading performance. However，when the proportion of dissipated energy is greater than 0.33，the initial unloading damage has significant effect on its reloading performance. (3) During the unloading，there is a negative correlation between the proportion of elastic energy and the proportion of dissipated energy，therefore in accordance with the correlation between the growth of dissipated energy during unloading and the damage development of rock samples，the ratio of dissipated energy during unloading and dissipated energy in unloading failure is proposed to represent the visual damage variable. (4) With the increase of unloading amount，the failure mode of damaged rock sample under reloading is gradually transitioned from typical shear failure to shear-tensile failure and tensile-shear failure. (5) Applying a lateral stress to the unloading damage rock sample can control the development of tension cracks to a certain extent，and thus improve the bearing capacity of rock sample.

The water weakening function is essential for long-term stability analysis and instability prediction of rock mass. Based on the mechanism analysis，a dynamic water weakening function to characterize the strength deterioration of dissolved carbonate rocks is proposed. X-Ray diffraction(XRD) of the rock mineral and scanning electron microscopy(SEM) experiments reveal that the calcite mineral content of Maokou Formation limestone is more than 99.6%. The acid etching can hardly change the proportion of material components but can aggravate the growth of dissolution pores，holes，and gaps and weaken the strength of limestone. The uniaxial compressive strength and dynamic softening coefficient were obtained from mechanical tests. The uniaxial compressive strength of the acid-etched limestone decreased between 50.8% and 57.2%，and its deterioration degree was higher than that of other stress states. The dynamic softening coefficient decreases with the increased etching time，which negatively affects the long-term stability of rock mass. The acid erosion rate，free acid absorption rate，saturated water absorption rate，and dynamic saturation were obtained from the water physical properties experiment，which revealed the influence of acidity and acid etching time on the difference in limestone dissolution. Acidity plays a primary role，but acid etching time determines the deterioration process of limestone in a weak acid environment. The decrease gradient of the dynamic water weakening function in the strong acid environment(pH = 2) is significantly larger than that in the weak acid environment(pH = 6)，and the maximum error between the predicted data of the fitting formula and the experimental results is less than 10%. The research results can provide critical support for studying dissolution rock mass deterioration and preventing and controlling geological disasters.

In order to analyze the influence of normal stress and rock strength on anchor bolt anchoring mechanism，three representative rocks，namely red sandstone，limestone and basalt，were selected to carry out direct shear tests of jointed rock mass under different normal stress levels. Based on the test results，according to the statically determinate beam theory and the principle of minimum complementary energy，this paper establishes a shear strength model of joint surface which considers the effect of rock structure and equivalent shear area. And combined with the theoretical calculation formula，this paper analyzes the influence law of rock strength on the axial force and shear force under different anchorage angles. The research shows the anchoring effect of anchor bolts is shown as increasing the equivalent internal friction angle and cohesion of joint surface，and that the greater the rock strength，the greater the contribution of the bolt to the shear strength of the joint plane. Compared with the experimental data and the existing analytical model，the calculated values of the theoretical model are more consistent with the experimental results. Through calculation and analysis，it can be seen that with the increase of the anchorage angle，the contribution of the axial force of the bolt to the shear strength of the joint surface gradually decreases，while the contribution of the shear force to the shear strength of the joint surface gradually increases，and the optimal anchorage angle of the bolt is within the range of 50°to 70°.

Due to the large vertical drop，the alpine glacier geomorphology is affected by vertical climatic conditions and local hydrothermal environment，resulting in typical vertical zoning characteristics. As an external dynamic form of dynamic equilibrium instability of epigenetic lithosphere，geological hazards must be closely related to its endowment of tectonic geomorphology and climatic environment. Based on the geomorphological features and statistical analysis of geological disasters in the glacial geomorphology of the Tibetan Plateau，this paper combed the typical characteristics of vertical zonation and endowment of each geomorphological unit，and then constructed the spatial allocation relationship between vertical zonation and geological disasters in alpine glacial geomorphology. The main results are as follows：(1) Alpine glacial geomorphology can be divided according to“four lines”，such as snow line，ice line，rain line and alpine tundra line，and there is a significant difference in boundary height between continental and marine glacial geomorphology. (2) The planation surface is important provenance of geological hazards in alpine glacial geomorphology area，which is obviously different due to the difference of glacier type. Specifically，there are two-grade planation plane provenance areas on mountain top and plateau surface in continental glacier area，while there are three-grade planation plane provenance areas on mountain top，Piedmont erosion surface and river valley erosion surface in marine glacier area. (3) There are typical vertical zoning characteristics of geological disasters in alpine glacial geomorphology according to the boundary mark of “four lines”. According to the boundary of “four lines”，the mode of dynamic geological action and altitude can be divided into：ultra-high ice erosion type(above snow line/ice-rock interaction zone). There are four types: high erosion type(between snow line and ice line/ice-water-cuttings interaction zone)，middle freeze-thaw denudation type(near ice line/ice-water-soil interaction zone) and low precipitation erosion type(below the alpine tundra line/water-soil-biological interaction zone). (4) There is a certain chain amplification effect of geological disasters between different boundary zones. the representative chain assemblage relationships are as follows：① ice(snow) rock avalanche→glocial lake and outburst flood→strong replenishment type mud flow；② ice(snow) rock avalanche→ice and snow debris flow→barrier lake and mountain torrent；③ moraine accumulation landslide→ice and snow debris flow→strong replenishment mud flow. The study can provide a basic reference for the regional and zonal prevention and control of major engineering geological disasters such as Belt and Road Initiative and Sichuan-Tibet Railway.

At first，for the smooth blasting of tunnel under high in-situ stress，the stress field between smooth holes induced by blasting was calculated by using the cylindrical cavity excitation model，and the influence of in-situ stress on the superposition of blasting stress between holes was analyzed. Then，the verified RHT numerical model was employed to simulate the circumferential tensile stress field induced by blasting under the conditions of single hole，double holes，three holes and in-situ stress，and the effect of radial uncoupling coefficient，hole spacing and in-situ stress on the cracking characteristic between smooth blasting holes was studied. Finally，some suggestions on the selection of smooth blasting parameters considering the influence of in-situ stress were given，and an application in blasting excavation of diversion tunnel at Jinping II hydropower station was presented. The results indicate that the superposition of blasting stress waves leads to the increase of the circumferential tensile stress near the midpoint of the line between smooth blasting holes，showing the characteristic of peak shaped bulge，while the radial compressive stress produced by the redistribution of in-situ stress is opposite to the action direction of the circumferential tensile stress induced by blasting between smooth holes，which will inhibit the crack formation and penetration between smooth holes；Influenced by in-situ stress，the propagation of blasting cracks shows obvious directionality and tends to expand along the direction of the maximum principal stress. Therefore，for the general smooth blasting design with the hole diameter of 42 mm，the charge diameter of 32 mm，the smooth blasting layer thickness of 0.6 m，when the in-situ stress increases from 20 MPa to 40 MPa，the hole spacing should be reduced from 0.75 m to 0.60 m，and the holes should be arranged along the direction of the maximum principal stress to ensure the formation of cracks between holes and shaping of excavation wall.

Coal-rock bursting liability is the natural attribute and the key influencing factor of coal-rock rockburst disasters. The coal-rock combined body was taken as the research object to accurately judge the bursting liability of coal and rock，and the uniaxial cyclic loading and unloading test was carried out on the coal-rock combined body，The elastic strain energy of the combined body under different stress levels was obtained，and the function between the elastic strain energy and the stress level was established. A new method for calculating elastic strain energy at peak stress time was proposed. A residual energy release rate index that comprehensively considers the peak strength，elastic strain energy，energy dissipation of the failure process and failure time of the specimen was proposed based on this，and the bursting liability judgment interval was given in combination with the existing indicators，and the rationality was verified. The results show that：(1) With the increase of the stress，the elastic strain energy shows a “slow-fast-slow” growth law，which corresponds to the compaction stage，elastic stage and plastic stage of the stress-strain curve. (2) The evolution laws of input energy density，elastic strain energy and dissipated energy density are similar to the evolution law of stress，all increasing with the increase of the stress. The increase of input energy density is the largest，and the increase of dissipated energy density is the smallest. (3) The elastic strain energy of the combined body at different stress levels was obtained through the experiment，and the functional relationship between the elastic strain energy and the stress level was established，that is，the square of the stress at any time had a good linear relationship with the elastic strain energy. A new method for calculating elastic strain energy at peak stress time was proposed. (4) Considering the peak strength of the specimen，elastic strain energy，energy dissipation in the failure process and failure time and other factors，a new bursting liability identification index was proposed：the residual energy release rate index WT，which is the ratio of the residual energy obtained by subtracting the post-peak dissipative energy density from the pre-peak elastic strain energy and the dynamic failure time represents the energy release in unit time when the coal-rock is impacted and damaged. This index is closely related to compressive strength Rc，bursting energy index KE，elastic energy index WET，and dynamic failure time DT. The evaluation interval of the residual energy release rate index WT was given：when WT≤0，the coal rock has no bursting liability；when 0＜WT≤14.5，the coal rock has a weak bursting liability；when WT＞14.5，the coal rock has a strong bursting liability. (5) The multi-sample method was used to verify the rationality of the proposed residual energy release rate index WT. This index was used to evaluate the bursting liability of coal seams in 8 mining areas，the results show that the evaluation accuracy of the residual energy release rate index WT is as high as 93.75%. The proposed new index to evaluate the bursting liability of coal rock has high accuracy and good rationality，which can reflect the actual bursting liability of coal and rock，and has good promotion and application value.

To reveal instant fracture of coal and abnormal gas emission at the disturbance of impact loads，the Hopkinson pressure bar test system of gas-bearing coal was used to carry out the tests about impact failure of three-dimensional gas-bearing coal and gas emission subjected to impact loads. The fracture evolution of raw coal samples and gas emission laws under instantaneous disturbance of impact load were studied，and the effects of loaded factors such as axial static load，confining pressure，gas pressure and impact load speed to the initial maximum gas emission speed were analyzed. The results showed that the three-dimensional gas-bearing coal presented fragmentation failure under the condition of low confining pressure or high impact load，while the coal samples presented axial tensile failure mode under other test conditions，and the size and number of cracks varied with the axial static load，confining pressure，gas pressure and impact load. The gas emission speed of raw coal samples reached the maximum at the moment of impact load，and then decayed in the form of power exponent. The duration of gas emission was basically within 2.00 s，which varied slightly with the change of loaded factors. The maximum gas emission speed at the moment of impact increased exponentially with axial static load，decreased linearly with confining pressure，and increased linearly with gas pressure or impact load velocity. Through the contribution of the loading factor change in unit scale to the maximum gas emission speed，it was determined that the importance of loading conditions affecting gas emission was r (impact load)＞r (confining pressure)＞r (axial static load)＞r(gas pressure). Based on gas emission model of coal particle expressed by Uskinov，the gas emission model of raw coal under impact disturbance was established. By the comparison of theoretical and experimental values about gas emission speed at different impact load velocity，their change trends and coincidence degree were good. This phenomena indicated that this model could reflect the effects of impact load to gas emission of raw coal samples，and be used to predict abnormal gas emission caused by impact disturbance in the stope. These achievements will help to reveal the formation mechanism about compounded dynamic disaster of gas-bearing coal，and promote the treatment of abnormal gas emission caused by stope disturbance，which will ensure deep coal resources mining in safe，efficient and green way.

In order to solve the problem of selecting the parameters of anchoring support in coal mine caverns，a method for selecting support parameters based on the distribution characteristics of surrounding rock plastic zone under different safety factors is proposed. Through the discrete element numerical simulation technology，the cohesion and internal friction angle in the parameters of rock block and structural plane are assigned and calculated according to the reduction coefficient. The range of surrounding rock instability region under different safety factors is obtained. According to the support and the actual requirements of the engineering，the relationship between the position of the anchorage section and the safety factor is analyzed，to achieve collaborative and quantitatively support parameters. The results show that according to the strength reduction method，the risk areas of complex caverns such as roadway intersections can be effectively analyzed，and the distribution characteristics of the safety factor of surrounding rocks can be obtained. In the support design scheme，the anchorage section of the short bolt should be ensured in the area with a safety factor greater than 1. For important parts such as the top of the roadway and the waist of the two sides，the anchoring end of the long anchor cable is set in the area with a safety factor greater than 2. The middle partition wall adopts double control anchor cables，and its reinforcement range is the area with a safety factor less than 1.3. Under the reinforcement of cooperative support，the measured maximum deformation value of the surrounding rock of the intersection tunnel is 2.9 cm，the maximum deformation value of the rectangular mining roadway is 27.5 cm，and the maximum error between the numerical calculation result and the field measurement result is only 0.5 cm，which shows that the support design method is highly reliable，and provides a scientific basis for the quantitative design of the support parameters of the coal mine tunnel.

TBM method is a mechanized excavation process applied to roadway and tunnel construction. In order to realize the optimization of operating parameters by automatic execution of TBM in the process of stable tunneling，an autonomous decision-making model of TBM control parameters based on the transfer learning method is established. First，the core optimization strategy is proposed，which aims to improve the tunneling efficiency and reduce the energy consumption on the basis of fully reference to manual experience. Taking the penetration rate per revolution(PRev) and the cutterhead revolutions per minutes(RPM) as output parameters，the mathematical modeling is completed. Secondly，an optimization model is constructed. Deep artificial neural networks is used as the basic architecture，including two auxiliary networks in the source domain and one main network in the target domain，which respectively performs the regression of control parameters，the prediction of rock breaking specific energy，and the optimization of control parameters. The methods of transfer learning and frozen-layers are adopted to achieve the unification of the source and target domain. Third，key hyperparameters are identified. The optimal hyperparameters of auxiliary networks are determined by a combination of orthogonal experiments and Bayesian optimization. The key weights of the objective function of the main network in the target domain are determined based on analytic hierarchy process. Finally，relying on the TBM construction data set(4 459 effective data) collected from the transferring water project from Songhua River in Jilin，the established model is trained and tested. The results show that in stable phase，the heading efficiency is increased by 15.55% on average，the energy consumption is decreased by 7.13% on average，and the variance is reduced，which improves the overall stability. The research results can provide technical support for the establishment of TBM intelligent control system.

In order to solve the problem of asymmetric floor heave of roadway and the potential of inducing kinetic events in dynamic pressure roadways，this paper established a“three-hinged arch-spring support”model based on the geometry and force characteristics of asymmetric roadway deformation，and analyzed the energy dissipation characteristics of the system and the sufficient discriminatory conditions for abrupt destabilization at the cusp of the total potential energy function of the system. The research showed that the deformation of the equilibrium curve takes two forms when along path I，the deformation is linear and progressive，and no large energy release kinetic event occurs. When along path II，the deformation is a non-linear and abrupt floor heave of roadway，where the lithology，dimensions and lateral stress of the roadway bottom are the key factors affecting the system instability，and the floor slab kinetic event can be prevented by pressure relief. Therefore，the technique of cutting the roof to unload the floor heave of roadway is proposed，the mechanism of pressure unloading is analyzed，and pressure unloading tests are carried out in the Caojiatan coal mine. The results prove that cutting the roof is done by changing the support form of the overhanging roof plate to shorten the collapse step of the roof plate，thus realizing the horizontal and vertical pressure unloading and the release of the collected energy in the quarry，and the horizontal pressure unloading is more obvious than the vertical，thus effectively controlling the deformation of roadway floor.

The study of surrounding rock failure mode is of great significance to the safe construction and efficient support of the underground caverns. The underground powerhouse is mostly constructed by layered excavation，and in this process，the stress state of the surrounding rock is constantly adjusted with the excavation，so it is necessary to study the surrounding rock failure mode of underground powerhouses with stress adjustment during the construction. Firstly，this paper summarizes the failure mode classification of surrounding rock in medium and low geostress levels by extensively investigating the surrounding rock failure mode in existing underground engineering. Then，the stress adjustment characteristics of the surrounding rock under layered excavation of shallow-buried underground powerhouses are analyzed. The failure mode classification of shallow-buried underground powerhouses considering excavation stage，typical parts，and rock mass basic quality(BQ) is proposed from the perspective of stress adjustment，and the phenomenon and mechanism of each failure mode are elaborated on and analyzed in detail. Finally，the failure mode classification is applied to the underground cavern group of Tuoba hydropower station. The failure mode classification of surrounding rock in this paper can provide some references for the prevention of the surrounding rock failure during the construction of shallow-buried underground engineering.

Exploring the evolution of pore structure inside rocks under high-temperature cycles and its influence on physical and mechanical properties is essential for long-term stability analysis of underground engineering，such as geological disposal of nuclear waste and geothermal development. In order to quantitatively analyze the effects of high-temperature cycling on the pore structure and physical and mechanical properties of granite，the evolution of surface characteristics，mass，volume，P-wave velocity，tensile strength，porosity，pore size distribution，and microstructure of granite under high-temperature cycling from 25 ℃ to 800 ℃ were investigated by a combination of nuclear magnetic resonance(NMR)，scanning electron microscopy(SEM) and thermogravimetric-differential thermal analysis(TG-DTA). The results show that：(1) With the increase of the temperature，the surface cracks，color difference，mass loss rate(Rm)，volume expansion rate (RV)，P-wave velocity attenuation rate(RP) of granite keep increasing，the tensile strength(σt) gradually decreases，and the physical and mechanical parameters of granite change significantly when T＞500 ℃. After five thermal cycles，the variations of physical and mechanical parameters of the rock are more prominent. (2) High temperature can promote the development of granite pores，and micropores and smallpores inside the rock gradually grow and connect to form mesopores and macropores，resulting in the enhancement of rock pore connectivity. Thermal cycling will further increase the connectivity between pore structures，leading to an increase in the proportion of mesopores and macropores and a further increase in porosity. (3) The deterioration of granite's physical and mechanical properties under high-temperature cycling is closely related to the changes in its internal pore structure. The Rm and RV increase linearly with the increase of the equivalent average radius of the pore(rT)，and the RP and σt increase exponentially with the increase of rT. (4) High-temperature granite will be water evaporation，quartz phase transformation，mineral oxidation，chemical bond fracture and other physicochemical reactions，resulting in intergranular cracks，transgranular cracks and peeling defects of mineral particles. Especially，the rock structure damage is more significant after quartz phase transformation at 573 ℃. (5) Thermal cycling will cause fatigue damage to the rock，and the alternating thermal stress generated by high-temperature cycling promotes the increase and expansion of microcracks，resulting in increased granite damage.

In order to explore the working characteristics of a new type of anti-floating structure in rock layer，that is，the ox leg anti-floating structure，based on the actual boundary conditions of the structure，the shear test device was designed and the indoor shear test was carried out to analyse the shear failure characteristics. The deformation evolution process，shear failure mode and damage evolution law were investigated by digital image correlation method and acoustic emission method. The results show that：(1) According to the characteristics of strength and deformation，the structural shear process can be divided into initial compaction stage，approximate linear elastic deformation stage，initial microcrack initiation and propagation stage，slow compression shear nonlinear deformation stage，stress brittle drop stage and plastic flow deformation stage. (2) The peak shear stress τp and peak shear displacement up show a“parabolic”trend of first increasing and then decreasing with the decrease of the leg angle，and show an approximately“inverse proportional function”trend of increasing with the increase of the leg height. (3) The failure starts from one fracture-prone area，and then successively breaks through in multiple fracture-prone areas. The shear failure mode is mainly characterized by compression shear on the left side，tension shear on the right side，and slip on the interface，with local small cracks. (4) With the increase of the leg angle，the failure mode changes from the slip along the interface to the penetrating failure along the interior of the material. With the increase of the leg height，the failure mode changes from the rapid pulling at the upper right corner to the slip along the interface and the reverse cracking at the leg to the upper and lower sides. (5) As the leg angle decreases，the duration of AE phase I is almost the same，while the duration of AE phase II increases first and then decreases，and the AE activity increases first and then decreases. As the leg angle increases，the duration of AE phase I is almost the same，and the duration of AE phase II increases gradually，and the AE activity increases gradually. The conclusion is helpful to guide the design of the ox leg anti-floating structure from limit state.

Due to the brittleness of rock and irregular joint，block-flexure is the most common type of toppling failure in anti-dip bedding rock slopes. Firstly，the failure mechanism of block-flexure toppling was clarified by centrifugal model test. A mechanical analysis model of block-flexure toppling was established. Then，based on the limit equilibrium method，the mechanical analytical formulas for the stability of continuous column and blocky column were derived respectively. The failure surface of block-flexure toppling was searched by the search algorithm. Finally，the whole analysis process was realized by MATLAB programming. The results show that the slope subjected to block-flexure toppling failure can be divided into four subzones(a toppling failure zone，a crack zone，a deformation zone and an unaffected zone). The failure surface of block-flexure toppling is stepped，and the height of these steps is equal to the multiple of the spacing of cross joints. It is found that the results of the theoretical analysis method are mutually verified with the results of centrifugal test. Then，the influencing factors were analyzed by the theoretical method. It is found that the inclination of cross-joint and inclination of slope face have a great influence on the scope of the first toppling failure zone；the thickness of column，tensile strength and the inclination of cross-joint have a great influence on the critical instability height. The research results can provide theoretical support for disaster prevention and control of the slope.

The dynamic change of karst groundwater level induced by heavy rainfall has an important effect on large landslides in karst mountainous areas in southwest China. In order to clarify the mechanism of karst landslides induced by heavy rainfall，taking Guanling landslide as an example，combined with field investigation，karst development phenomenon and seepage field characteristics，this paper analyzes the genesis mechanism of Guanling landslide，and finds that，under the condition of heavy rainfall，the strong lateral recharge at the back of the slide source area raises the groundwater level，which causes the formation of high pressure water head in the slope body of the fissure water area of bedrock and induces the landslide instability on the potential weak structure plane in the slide source area. Then，the stability of the sliding body under multiple working conditions is analyzed by means of groundwater table analytic decoupling residual thrust method. The theoretical calculation results show that the average diameter of karst pipeline medium and rainfall intensity have great influence on the stability of the slope.  When the average diameter of karst pipeline is 0.15 m or above，slope instability occurs under extreme rainfall conditions. The calculation formula of groundwater level and stability analysis method derived in this paper can provide useful reference for stability evaluation and prevention of karst landslides.

Deep buried central gutter can effectively reduce the high water pressure in the water-rich area of the elevated arch and control the uplift of the tunnel bottom caused by high water pressure. By designing a 3D printed tunnel model test，the influence of the important design parameters of the deep buried gutter on the drainage efficiency and structural safety of the tunnel is investigated，and a reasonable range of values for the important design parameters of the deep buried gutter is given. The results of the study show that: the deep buried gutter pipe diameter parameter is the dominant factor affecting the tunnel drainage，lining water pressure and structural displacement，and there are phases in the pattern of influence，in 0–0.8 m pipe diameter conditions，with the increase of the pipe diameter，deep buried gutter drainage increased significantly，the distribution of lining water pressure gradually changed from“scallop type”distribution to“peach type”distribution，and the bottom structure gradually changed from an overall significant uplift state to a slight subsidence state. In 0.8–1.6 m pipe diameter conditions，with the increase of the pipe diameter，the total tunnel drainage，lining water pressure distribution and bottom structure displacement maintained a steady state of less variation，taking into account other control factors，the recommended value of pipe diameter in similar projects is 0.6–1.0 m. The burial depth parameter is not the dominant factor affecting the drainage efficiency，but should be selected to facilitate site construction，ensure the drainage effect and contribute to the stability of the tunnel bottom，etc.

Aiming at the problem of negative friction of pile foundation in collapsible loess area，a kind of crawler steel pipe pile is proposed，which can not only reduce the negative friction of pile side，but also increase the positive friction. Its structure and working principle are discussed. Based on the load transfer function and combined with the motion characteristics of the track，the load transfer method is modified according to the motion characteristics of the track，and the distribution functions of the positive and negative friction on the side of the pile，as well as the calculation formulas of the pile settlement and the soil settlement considering the time effect are derived，Finally，the calculation method of the vertical ultimate bearing capacity of the new pile is given. Combined with an example，the structural characteristics and working principle of the new pile are simulated by using the nonlinear finite element software ABAQUS，and the simulated value is compared with the theoretical value. The indoor model test of new type pile and common pile is carried out. The results show that：(1) The track can make the neutral point move up and reduce the negative friction on the pile side. (2) With the increase of the distance from the pile center，the smaller the settlement is，which is more conducive to the stability of the foundation. (3) The simulation results of single pile bearing capacity are basically consistent with the theoretical calculation results，which verifies the correctness of the theoretical analysis method. (4) Compared with the ordinary straight pile，the effect of reducing negative friction is obvious，and the bearing capacity of single pile is improved，which provides certain theoretical guidance for the design of pile foundation projects in collapsible loess areas.

The movement patterns of retaining wall has important influences on the distribution of earth pressure and the soil deformation and failure modes of behind the wall. Using the developed valuable retaining wall model test apparatus and quartz sand filler，model tests were carried out in five patterns：translation，rotation around the bottom，rotation around the top，counter clockwise rotation around the midpoint point and clockwise rotation around the midpoint point. The measured earth pressure was compared with the existing calculation solutions，and the earth pressure at rest，ultimate earth pressure and the variation law of earth pressure with deformation were analysed. At the same time，the soil deformation characteristics and failure modes of soil were analysed using the particle image velocimetry technology. The results show that the earth pressure at rest increased linearly along the wall depth，while the measured value was greater than the theoretical solutions due to the over consolidation of sand. There are great differences for the displacement required for reaching the limit state under different displacement patterns. Due to the high compactness of the sand，the measured value of the limit earth pressure of translation，rotation around the bottom and rotation around the top was less than the theoretical solution. The action point of resultant force of the limit earth pressure of counter clockwise rotation and clockwise rotation around the midpoint moved up and down respectively compared with that of rotation around the top and bottom. The slip shear planes of different displacement patterns showed different shapes，and the inclination angles of the measured slip planes was slightly larger than the theoretical values. Rotation around the bottom and clockwise rotation around the midpoint showed progressive failure mode，while translation，rotation around the top and counter clockwise rotation around the midpoint show integral failure mode.

Dynamic shear modulus and damping ratio are significant dynamic indicators of soil dynamics and geotechnical earthquake engineering，and their constitutive relationships are mostly established and recognized based on unit tests. Compared with the unit tests，the centrifuge model tests have more realistic stress boundaries，loading schemes and drainage conditions，but there are rarely studies on dynamic modulus and damping ratio. The form differences of the shear stress- strain hysteresis loop of dry sand and saturated sand are revealed，and the characteristics and laws of dynamic modulus and damping ratio with burial depth，shear strain and vibration order are compared based on the centrifuge model comparison tests. The results show that the hysteresis cycle of dry sand is oval，but the hysteresis cycle of saturated sand is irregular and non-smooth dumbbell shape. The shear strain of saturated sand increases rapidly after liquefaction，and the maximum shear strains of dry sand and saturated sand under 0.2 g load are 0.3% and 1%，respectively. The dynamic shear moduli of dry sand and saturated sand obey the Hardin hyperbola model well，which decrease with the increase of shear strain，increase with the increase of burial depth，and increase with the increase of vibration times，which causes the soil vibration density. The maximum dynamic shear modulus and their changes with depth and vibration order are basically the same，and are less affected by saturation. The reference strains of dry sand and saturated sand are 0.1%–0.35% and 0.05%–0.18%，respectively，and the reference strains of dry sand under different depths and loads are about twice that of the saturated sand，and they increase with the increase of burial depth and density. The damping ratio of dry sand has a good law with the variation of shear strain and burial depth，and the discreteness is small. However，the damping ratio of saturated sand is not obvious，and the discreteness is large，which is related to the change of sensor position，the test accuracy and the reliability of the analysis method caused by the failure of the soil strength after liquefaction. The research results provide technical support and guidance for understanding the response characteristics of the dynamic modulus and damping ratio of the centrifuge model test and for comparing the reliability of the dynamic modulus and damping ratio in the unit test.