ISRM International Symposium - 6th Asian Rock Mechanics Symposium

ABSTRACT:

Underground haulage drifts serve a crucial purpose of moving blasted ore material from the stopes. These drifts are driven in either footwall rock or hanging wall rock depending on the geology of the mine. As there is constant movement of men and material, the stability of these drifts is of critical importance and every measure is taken to make sure that these drifts are stable and functional. Significant effort is put into understanding the stability of these drifts and reinforcing them with suitable supports during the mine life. By making use of some user defined and built-in functions available in FLAC, a simple method to implement multiple simulations using random variables has been developed to determine the stability of underground haulage drift in the vicinity of mining activities, using numerical modelling. The probability of unsatisfactory drift performance is derived based on pre defined performance function of drift wall deformations.



1. INTRODUCTION

Numerical modelling has become increasingly popular as a tool for geomechanics mine design and mine stability analysis. Methods like the finite elements, boundary elements and finite difference have proven to be extremely useful, and have been successfully used for the development of a wide range of commercially available numerical modelling software. In practice, commercial numerical modelling software is often used to perform deterministic analyses and the design values are selected accordingly. For example, the material is considered to be elastic or elastoplastic and the stability is assessed by examining the extent of yield zones, the deformations causing drift wall convergence or roof sag, etc. and, the design of support system for underground openings is based on a combination of past experience, empirical methods, and deterministic numerical models. Because of the deterministic nature of numerical models, model parametric studies or sensitivity analyses, in which model input data are varied within plausible range, are often carried out to allow for better understanding of the problem before hand, e.g. stability of mine openings, as a result of constant variation in estimated rock mass properties. Although, deterministic numerical modelling has been quite useful in estimating failure, it has been well recognized by many practitioners that numerical modelling alone cannot predict the probability of failure. Because of the inherent uncertainty associated with parameters like the in-situ stress fields, rock properties and geological features around the openings, there is also high uncertainty in the selection of support design based on such parameters. Thus, there is a need to perform stochastic analyses to derive a probability of failure to better understand the risk associated with choosing design parameters based on uncertain input data. Stochastic analysis techniques have been used with success in conjunction with limit equilibrium methods (LEM) to calculate the probability of failure in geotechnical applications like slope stability and isolated pillar collapse (Fenton and Griffiths, 2008). These problems are characterized by a single equation defining the safety factor as a function of one or two random variables, typically cohesion and angle of internal friction of the rock.

Abstract:

The understanding of time dependent effects or creep behavior is important in further development of knowledge in the field of rock mechanics. Increasing pressure on support system because of creep behavior of rock is one of the most important problems in underground structure which surrounded with weak rock mass. This paper presents a time dependent behavior analysis of Siah Bisheh pumped storage powerhouse cavern with complex geometry, available geological formations and diverse geotechnical properties of rocks, is under construction on the Chalus River at the north of Iran. Because of the fractured and jointed rock mass, the Discrete Element method Code 3DEC used to compute convergence and pressure on the support system. The power constitutive creep model which is able to model the primary and secondary creep regions of rock masses is applied. In this paper, results of the Siah Bisheh cavern as a case study dealing with a comparison between long-term deformation data from in situ measurements and obtained data from an analytical solution and numerical simulations with special consideration of anisotropic effects will be presented. The differences between the obtained values might be because of discontinuous media or different inputs parameters were used for analysis.



1. Introduction

The rheology and time-dependent behavior of rock are fundamental mechanical properties of the rock material, which can be used as an important bases in explaining and analyzing the long-term stability for rock engineering. Especially, time dependent deformation of rocks has significant effect on stability of underground structures, such as nuclear waste storage facilities, tunnels and powerhouse caverns. It is particularly common in underground structures excavated in soft rock, heavily sheared weak rock masses or rock masses subjected to high in-situ stresses (Verman, et al., 1988; Barla, 1995; Bhasin & Grimstad, 1996). In order to study the stability of the underground structures and designing their support system, time dependent deformations should be highly considered (Shalabi, 2004; Tsai, 2008). Therefore, time dependent behavior of underground structures as well as predicting the long-term behavior, is of great importance. This paper presents a time dependent behavior analysis of Siah Bisheh pumped storage powerhouse cavern with complex geometry. The cavern is being built in a region that is highly prone to sheared and fault zones where their failures could cause severe damages to underground spaces. It is therefore essential to analyze and design underground structures for preventing any serious long-term damages in the region. The rock mass may exhibit continuous or discontinuous deformations due to the existence of large underground openings: shear of joints and creep deformation.



2. Siah Bisheh pumped storage project

The Siah Bisheh pumped storage project is located at the Alborz chain Mountains being mainly formed and folded during the Alpine orogenic phase. The most important tectonic phenomenon of Siah Bisheh area is the fault called as the Main Thrust Fault (MTF), with a dip direction of 78/028 and an almost E-W trend. The site of powerhouse caverns is generally located at the Permian Formation part.

Abstract:

The application of the numerical Discontinuous Deformation Analysis method (DDA) in rock engineering is discussed here. Following a review of recent 2D and 3D DDA validations, application of the method in analysis of natural rock slopes and underground openings are presented. Two interesting problems are explored in the applied section of this paper: (1) the deformation of discontinuous overhanging rock slopes, and (2) the stability of karstic caverns underneath active open pit mines. The numerical DDA method is shown here to be a powerful tool for modeling dynamic rock mass deformation when the interaction between multiple discrete elements dictates the expected global deformation.



1. introduction

Recent developments in the validation and application of Discontinuous Deformation Analysis (DDA), originally developed by Shi and Goodman [1] are presented. We begin with a brief review of recently published 2D - DDA validations for cases of dynamic loading [2] along with new 3D - DDA validations for single and double face sliding [3]. Following these validations we present dynamic DDA applications in natural rock slopes and underground openings.



2. Dynamic DDA Validations

2.1 Single Face Sliding in 2D

A displacement based sliding block model was first proposed by Newmark [4] and Goodman and Seed [5], is now largely referred to as “Newmark” type analysis. Determination of the amount of displacement during an earthquake involves two steps [5]: 1) Determination of horizontal acceleration required to initiate down slope motion, also known as “yield acceleration” (ay), which can be found by pseudo-static analysis, and (2) Evaluation of the displacement developed during time intervals when yield acceleration is exceeded, by double-integration of the acceleration time-history, with the yield acceleration used as reference datum.



2.2 Single Face Sliding in 3D

The two dimensional formulation shown in Equation 4 has been expanded by Bakun Mazor [3]to three dimensions for a single block on an inclined plane, in order to validate 3D-DDA under three components of input motion. The analytical solution is based on the original static limit equilibrium formulation presented by Goodman and Shi [6]. A typical three dimensional model of a block on an incline is illustrated in Fig. 2(a). The dip and dip direction angles are, α = 20˚ and β = 90˚, respectively. A Cartesian coordinate system (x,y,z) is defined where X is horizontal and points to east, Y is horizontal and points to north, and Z is vertical and points upward. In an unpublished report, Shi [7] refers only to the case of a block subjected to gravitational load, where the block velocity and the driving force have always the same sign. The same is true for the original equations published in the block theory text by Goodman and Shi [6]. For the complete set of equations of the three dimensional solution see [3]. The relative error of the new analytical solution and 3D-DDA method with respect to the existing Newmark's solution is shown in the lower panel of Figure 3(A) and are both found to be less the 3% in the final position.

Abstract:

The paper reports the described methodology applied to a case study of stability evaluation of a rock slope at high elevation (Monte Bianco) where instability phenomena where observed. Stability analysis were performed using both Limit Equilibrium Method (LEM) and Distinct Element Method (DEM) in 3Dimensional field based on accurate non contact surveys. The surveys are supported and validated by traditional in situ measurements of the discontinuity orientation and other rock mass features. The required discontinuity shear strength has been validated by direct shear tests performed in laboratory on specimens taken on the slope. Furthermore, the paper presents some considerations on the influence of climate changes on the rock mechanical features and examines the effect of temperature variation on the stability conditions of rock slopes; this is done by means of several parametrical analysis of the Aguille Marbrèe rock outcrop carried out modifying the rock discontinuity saturation, persistence and frequency, shear strength etc.. These analysis are showing how a different water regime in the rock mass, caused by climate variation, can strongly influence the stability condition of the slope; on the other hand, the DEM analysis are used to better understand the complex and progressive instability phenomenon that was observed on site. The comparison between the shape of the unstable blocks and their volume as computed by DEM and those measured on site at the base of the slope, validate the DEM results.



1 INTRODUCTION

In the Italian North West alpine region, several instability phenomena involving large rock volumes have been observed in the last years at high elevation (3000 m a.s.l.). The increment of these phenomena occurrences could be related with the registered climate variation and with a general tendency of temperature increment. The research presented in this paper was dedicated to the development of a methodology for a better understanding and forecasting of the rock slope stability condition at high elevations. The evaluation of rock slope stability is based on the geostructural characterization of the rock mass and, therefore, structural surveys are the basic requirement of any stability analysis. Since hiking and climbing are among the most important economic resources for alpine regions, the need to improve the safety condition at high elevation is of primary importance and, consequently, rock slope risk evaluation becomes more and more demanding in terms of objectivity and details. To become a real alternative to a traditional survey (both in terms of productivity and accuracy) the photogrammetric or laser scanner techniques requires interactive or automated software tools to allow the efficient selection of discontinuities or the interpretation of the normal vector pattern. For this reason, the segmentation algorithm of the point cloud in subsets, each made of points measured on a discontinuity plane of the rock face has been implemented in RockScan [1], a software tool developed to facilitate the interaction with the point cloud in the identification of the discontinuities; rather than on the 3D data, selection of regions of interest is performed on oriented images of the rock face.

SYNOPSIS

Post-failure deformation experiments were carried out for rocks and coal having compressive strength varying from 24 MPa to 366 MPa and analyzed for the two classes of post-failure behavior i.e.Class-1 and Class-II as reported in the literature. All the experiments were conducted under uniaxial compressive stress conditions using an MTS compression testing machine. Depending on the rock type, either lateral strain or displacement control was used as the feedback signal. Post-failure deformation was recorded for all the rocks and coal samples. Based on the post- failure deformation pattern and the fracture mode of the failed samples, three types of post-deformation patterns were identified. Type-1 is characterized by a steep fall in stress with tensile cracks, Type-2 showed a typical pattern of rise and fall in stress with a combination of tensile and shear cracks, while Type-3 showed a gradual decrease of stress with shear cracks. The Class-II failure as reported in the literature was not observed even for rocks having strength of 366 MPa. Based on this study it is inferred that all the rocks may undergo post-failure deformation, the amount of deformation depends on the failure mechanism which is a combination of tensile and shear cracks.



1. INTRODUCTION

The concept of Class-I and Class-II post-failure behavior was originally proposed by Wawersik and Fairhurst(1,2) to classify the shape of the complete stress-strain curve for a particular rock according to its strain beyond the peak strength. If the strain increases monotonically throughout the failure process, the curve is designated as Class-I, and other curves as Class-II. Class-I behavior is characterized by 'stable' fracture propagation. But rock that exhibit Class-II behavior, the failure is unstable and the fracture of rocks cannot be controlled. The dividing line between Class-I and Class- II behavior is defined by the dashed line as shown in figure 1. In the present study six different rocks including coal having strength from 24 MPa to 366 MPa were investigated for the two types of post-failure behavior.



2. SAMPLE PREPARATION

A total of six rock types including coal were investigated, table 1 gives their uniaxial compressive strength, Young's modulus and Poisson's ratio. Except for coal remaining samples were collected in the form of drilled cores. Coal samples were drilled from the blocks perpendicular to bedding plane to obtain NX size cylindrical cores. The diameter of the samples varied from 37 mm to 54 mm. The length to diameter ratio was maintained at 2.0. Samples were cut to the required length and the end faces were ground using surface grinder.



3. EXPERIMENTAL METHODOLOGY MTS

rock mechanics system was used along with their extensometers for conducting post-failure experiments. As reported in the literature (3, 4, 5, and 6) three feedback controls are commonly used for the post-failure tests and they are:

1. Lateral strain control

2. Axial strain control

3. Displacement control or also called as stroke control Load control is not used as a feed back as it produces sudden failure of the sample beyond the peak stress.

Abstract:

A good understanding of the subsurface distribution of rock properties and its uncertainties are important elements in the development of safe and economical geo-engineering projects. The aim of this paper is to describe the steps to be followed in order to construct a comprehensive model of the rock mass. A brief review of the main elements of 3D geological and geomechanical modeling is presented followed by a discussion of the main tools available for applications to rock engineering. One case history that describes the evaluation of rock mass conditions for a powerhouse excavation in a metamorphic environment is presented. Geostatistical techniques were used to construct a model of rock fracturing and rock mass strength throughout the building area. After the excavation the rock model was compared with conditions displayed by the actual rock mass.



1. INTRODUCTION

The study and understanding of a complex environment of difficult access, the subsurface, has always been a challenge and the objective of geoscientists and some related fields, like geotechnical, mining and petroleum engineering. Engineering works are designed and investments are secured based primarily upon site characterization. Therefore, good site characterization, with distribution of rock properties and its associated uncertainty, is of fundamental importance for the decision process and management of geo-engineering projects. Traditional geotechnical engineering works are usually based on 2D representations of data between boreholes (the so-called geotechnical cross-sections). These cross-sections are not able to fully represent the global distribution of properties on the subsurface though carried out by highly skillful geologists. Industries like oiland- gas and mining, on the other hand, use sophisticated 3D modeling techniques to predict subsurface property distribution and its associated uncertainties, a process called 3D geological characterization [1]. The geological characterization comprises actions for the acquisition and interpretation of data from different sources and scales combined with a 3D representation and hierarchical importance for developing a model. The modeling procedure was first used in the mining industry and later was adapted and improved by the oil-and-gas industry, which introduced the concept of Common Earth Model or Shared Earth Model. The construction of these models depends very much upon the existence of cross-disciplinary teams in order to unite the knowledge of different areas [2]. Following the examples of these industries, the geotechnical community has made some efforts to change the traditional practice. Recently, many studies that employ this technique to develop 3D models for site characterization have been published [3-7]. This process, however, has evolved slowly, because civil engineers argue that the costs to develop these models are too high [8]. An explanation about geological and geomechanical elements involved in the model build up process is presented and one case history is described to illustrate the proposed workflow.



2. METHODOLOGY

The first step during the development of a 3D earth model is to define the final objective, i.e., what are the models for? Who are they for? [9]. Once the final objective of the model has been defined, data requirements and their collection procedures can be established.

Abstract:

Research into metal cutting is an established area, whereas the behaviour of rock during cutting is somewhat more complex and not well understood. This makes the field of Rock Mechanics - in particular the use of Finite Element Analysis [FEA] in studying the behaviour of the rock under cutting conditions where many parameters are involved - a challenging area to explore. Here experimental results are needed to justify and verify the results from FEA modelling. Based on the Brazilian Test [1] this paper analyses three types of rocks namely - limestone, sandstone and shale, which belong to the sedimentary rock category with sandstone having greater strength than limestone and shale. The Finite Element-Discrete Element coupled method (FEM-DEM) used in the proposed modelling enables the facture process of the rock to be studied closely in relation to the cutting force component causing the start of fracture and the behaviour of the propagating crack during this process [2]. These results should prove to be useful in analysing the fundamental behaviour of the rock material at the extreme cutting edge of the tool.



INTRODUCTION

Rock being a natural material presents a high degree of complexity in understanding its response to a given state of stress. Rocks are made up of different elements (Si, Na, Ca, Al, Fe, O, C, etc) which constitute the minerals that are bonded by a cementing agent, usually silicon. The most common minerals are Quartz, Feldspar and Mica; The different mineral make up, the temperature and pressure at which rocks are created, all add to the inherent anisotropy present in any rock mass. Furthermore it also contains discontinuities such as faults and fractures and fluid-filled pores. Rocks tend to be anisotropic, hence a finite element analysis technique employed to model the mechanical characteristics of the rock should be able to take into account the above mentioned discontinuities [3]. The most commonly used cutting tool materials are Tungsten Carbide and Polycrystalline Diamond Compacts. They are used in machining of rocks in various industries and tasks such as drilling for oil and gas, extraction of geothermal energy, mining, excavation, tunnelling, coring and for rocks used as building materials. Rocks are extremely variable substance such that studies based on one particular type of rock may not be applicable to other rock types. Rock fracture mechanics studies the failure criterion and the failure mechanisms of rock fracture and provides ways to either maximise or minimise rock fracture [2]. Wear and tear of the tools is inevitable and thus there is a greater need to understand the fracture mechanics of rocks in order to improve tooling. Numerical modelling has been successfully used to model material removal process in rock cutting [4], to understand the influence of friction on chip size [5], to study the material deformation of heterogeneous rocks [6-8], study of rock bolts [9], to name just a few. An exhaustive review of numerical modelling related to rock mechanics is provided by Jing [3].

Abstract:

Indian coalfields inherit difficult geo-mining conditions, which make it difficult to extract developed coal pillars by an underground mining approach. However, to meet the demand of coal production, three attempts of application of the mechanised depillaring have been made with mixed results. This paper, first, briefly presents results of the country's first fully mechanised depillaring face and, also, discusses the outcomes of two other attempts of mechanised depillaring. A brief review of the exiting geo-mining conditions Indian coal-fields shows that the role of support, both, natural and applied support is an important issue, which need to be addressed for the varying conditions of different sites. The important technical point to notice during the first fully mechanised depillaring face is successful application of high capacity, pre-tensioned, stiff and resin grouted roof bolts as systematic support of roof (SSR) and breaker line support for the laminated roof of the site. A review of different mechanised depillaring faces shows that the high capacity roof bolts and mobile roof support are complementary for weak and moderate roof strata. But, this approach of roof support may not always be applicable in India due to frequent encounter of massive roof strata. It is an open fact that pillar recovery at higher depth cover is another important emerging issue in the country. Coal mining activity at deeper cover is likely to intensify in the future because the shallower deposits are exhausting. At greater depths, the size of pillars is too wide to be fully extracted with single pass pillar stripping techniques. Here pillar splitting before extraction becomes the only alternative if the pillars are to be fully extracted. On the basis of experiences of different conventional semi-mechanised depillaring faces, this paper also discusses the rock mechanics aspect of manner of pillar extraction and efficacy of the left out fender for a successful high speed depillaring. It is realised that the experience of local rock mass behaviour is an important input for a successful adoption this mass production technology.



Introduction

Future extraction of coal, mainly through underground mining, is going to encounter, relatively, deeper excavations in difficult rock mass conditions. In India, large number of coal seams has extensively been developed by formation of pillars to meet the increasing demand of coal in the country. This is found to be a simple and safe method of coal production from underground mines. The process of pillar formation also received favourable situation due to presence of the competent coal seam and similar surrounding rock mass. Indian Coal Mines Regulations (CMR), 1957 is also quite liberal for pillar formation. This strategy of coal production found suitable for the Indian coal mining industry because of low capital investment and involvement of trivial technical expertise. Now, industry is looking towards the huge amount of coal locked up in the pillars. Underground coal mining in India is facing serious techno-economical challenge during depillaring of the developed coal seams. Conventional depillaring frequently encounters strata control problem and the productivity remains quite low.

Abstract:

A large number of shotcrete slopes that were constructed in Japan during the 1970's are now more than 30 years old, and there has been a significant degree of deterioration. In Japan, many of the slopes covered with shotcrete have aged considerably. Therefore, there is a risk that slope failures may occur due to the effects of factors such as seasonal weather patterns, natural disasters, climate change, heavy rainfall, and earth quakes. Therefore, it is important to develop a method for monitoring the stability and the durability of these slopes. In this paper, we propose a technique that converts seismic velocity and electric resistivity data to porosity and saturation, which is then used to monitor weathering and groundwater fluctuation behind the slope. The evaluation results of this methodology confirm that the distribution of porosity and saturation of rock mass around the evaluated slope arrived at by this conversion system agree with those of the actual rock mass conditions evaluated using boring samples. In addition, it was possible to monitor the signs of seasonal variation and weathering in the ground by performing monitoring using this methodology over a period of multiple years.



1. INTRODUCTION

Slope stabilization structures made with shotcrete were constructed in large numbers in Japan in the 1970's. As more than 30 years have passed since their original construction, a variety of problems have come to light, such as cavitations behind the concrete lining, and slope failures caused by weathering of the natural ground. Currently slope management is conducted primarily by road patrols and road facility inspections to prepare disaster prevention records regarding aged slopes. If further investigation of a particular slope is found to be required, a geophysical exploration is conducted to review countermeasures to be taken against defects. For these reasons, Kusumi and Nakamura [1] devised a technique that can convert in-situ seismic velocity and resistivity to porosity and water saturation distributions (hereafter referred to as “conversion analysis”). In this study, using this conversion analysis methodology we tried to monitor the thickness of weathered rock and the degree of moisture build up in a more accurate manner, with a focus on seismic and electrical explorations that are considered to be effective in assessing the weathering of the natural ground behind the slope and variations in ground water levels.



2. GEOLOGICAL CONDITIONS OF THE SLOPE

The geophysical data for the analysis in this study was acquired at a shotcrete slope (A-district) and a non-supported slope (B-district) of a national highway as shown in Figure 1. In geological terms, the overall slope consists of a sandstone layer, alternating sandstone-shale strata, and a basalt lava layer, which are located in the Tanba strata formed during the Triassic and Jurassic periods of the Mesozoic era. Figures 2 and 3 show the locations of the investigation, which are near the south side of the national road. The overall slope is comparatively large-scale, with dimensions of about 200m in length and about 50m in height.

Abstract:

The McArthur River Operation is an underground uranium mine operated by Cameco Corporation with a joint venture partnership with Areva Resources Canada. Accessing the exceptionally rich orebody has been challenging due to a weak host rock mass at depth, pervasive fault and joint systems, a high level of radioactivity, and the water pressure resulting from the 500m-thick water-bearing sandstone formation. Mining and geotechnical engineers at McArthur River Operation have leveraged past experiences in the Athabasca Basin, artificial ground freezing, a hybrid of mining and civil engineering excavation techniques, and a novel integrative design and construction process to successfully access very challenging ore zones. The most outstanding achievement was witnessed during the design and construction of the first raisebore chamber in the Zone 2 Panel 5 area. The safe completion of this development was a 'game changer' for future development, mining sequencing and ore recovery.



INTRODUCTION

Located in the Athabasca Basin of Northern Saskatchewan, the McArthur River Operation is the world's largest high grade uranium mine, as it provides 15% of the world's uranium production with an average annual output of nearly 9,000 tonnes of uranium oxide in 2009 [1]. The mine hosts enormous reserves of high grade ore (approximately 21%) occurring between 500m and 640m below surface [2]. Controlled mostly by the P2 thrust fault, the unconformity-type orebodies at the McArthur River Operation are nestled between the crystalline basement rocks of the Aphebian Wollaston Domain, and the sandstones and conglomerates of the Helikian Athabasca Group [2]. Around the ore zones, the radioactive nature of the deposits has created a halo of altered rock masses at various stages; clays, pervasive clay infillings, and unconsolidated sands add to the complex fracture system. One solution to mitigate these challenges has been the use of the raisebore mining method to safely and economically extract the ore. Although innovative and effective, the method still requires raisebore chambers and extraction chambers to be excavated in complex geological and hydrogeological conditions. Geotechnical endeavours have always been carefully engineered at the McArthur River Operation to account for all the parameters. However, in the geotechnical design and execution of the new Zone 2 Panel 5 (Figure 1), a novel approach that integrated 10 years of experience in the Athabasca Basin was crafted and successfully implemented. This paper recounts a new geotechnical design procedure, construction, and monitoring that is considering by the author to be the best-practice for mining through the unconformity in the Athabasca Basin.



DESIGN

As described in previous papers [3-4], the McArthur River Operation has created and successfully implemented a novel geotechnical design procedure to account for geological and hydrogeological conditions, geotechnical analyses, radiation protection, artificial ground freezing, inflow preparedness, customized factors of safety, and regulatory requirements (Figure 2). The details about each of these aspects can be found in Yameogo [4]. In the case of the first chamber of the Zone 2 Panel 5, three geomechanical domains were identified with a specific excavation sequence and design, timing, and installation of the ground support [3].