STABILITY ANALYSIS OF A LONGWALL MINING IN NARVA OIL SHALE MINE

Transcription

1 STABILITY ANALYSIS OF A LONGWALL MINING IN NARVA OIL SHALE MINE Ott Oisalu Taavi Lõhmuste Civil Engineering, master's level (120 credits) 2017 Luleå University of Technology Department of Civil, Environmental and Natural Resources Engineering

2 LULEÅ UNIVERSITY OF TECHNOLOGY Civil engineering X7003B Taavi Lõhmuste, Ott Oisalu STABILITY ANALYSIS OF A LONGWALL MINING IN NARVA OIL SHALE MINE Master s degree project Supervisor: Dr. David Saiang External supervisor: Dr. Oleg Nikitin Luleå 2017

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4 TABLE OF CONTENTS Abstract... 5 Acknowledgement Introduction Background Project organisation Tasks of the project Methodology Baltic Basin oil shale resources Characterization of oil shale in Baltic basin Kukersite distribution in Estonia and geological strata Characterization of Estonia deposit Narva mining field backround information and location description Narva oil shale quarry geology Geotechnical model Description of soils and rocks Description of discontinuities General geomechanical properties of rock Geomechanical properties of rock Hoek & Brown parameters Model parameters rock mass properties General model description Hydrogeological description Model description for the highwall RS Model description for the highwall FLAC Technology of (punch) longwall mining Chain pillar design Empirical method Results Numerical modelling results for the highwall slope stability Results from ALPS (Chain pillar stability) Discussions and recommendations Higwall slope stability analysis Chain pillar stability analysis References Appendices Appendix 1. Geological crossection of Northern-Kiviõli oil shale quarry [36] Appendix 2. Production layer thickness along Estonia deposit [11] Appendix 3. Borehole data (see Figure 3) [18] Appendix 4. Production layer physical parameters [37] Appendix 5. Safety factor for a 3-entry system Appendix 6. Safety factor for a 2-entry system Appendix 7. Safety factor related to length of the pillar (undersized width)

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6 ABSTRACT STABILITY ANALYSIS OF A LONGWALL MINING IN NARVA OIL SHALE MINE Taavi Lõhmuste, Ott Oisalu Civil Engineering Luleå University of Technology Year 2017 Oil shale industry in Estonia is looking at other mining technologies as alternative to strip mining and room and pillar mining methods. One such alternative to the room and pillar method is the punch-longwall mining method. Enefit Kaevandused AS, one of the major oil shale companies in Estonia, plans to employ this technology in exploiting some of its resources in the near future. This thesis examines the different stability problems related to the planned punch-longwall mining project in Narva oil shale mine. Determining optimal chain pillar dimensions and stability of the punch-longwall highwall slope are the main objectives of this project. Rock mechanical analyses have been done and recommendations are made based on the rock mechanical aspect of the mining process. Taavi Lõhmuste is responsible for the chain pillar stability analysis and Ott Oisalu for the punch-longwall highwall slope stability analysis. It is essential to understand the geology of a certain area in order to make accurate stability assessments. Because of the previously stated requirements, the geology of Estonian oil shale deposit is examined in the first part of the thesis in order to determine the geological and rock mechanical conditions to set the foundation for further analyses. In conclusion, for the part of the highwall slope, a properly designed barrier pillar plays a key role in the stability of the slope. After reviewing and analyzing the results of both highwall slope numerical models, it can be stated that the minimum length for the barrier pillar that still will yield in stable highwall slope is 65 meters. For the part of the chain pillars, in conclusion, it can be determined that optimal chain pillar dimensions that should be suitable, from the stability standpoint, are 6x6 meters for 3-entry system and 7x7 meters for 2-entry system (length x width). Keywords: longwall mining, oil shale, stability analysis, chain pillars, barrier pillar, slope stability 5

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8 ACKNOWLEDGEMENT We would first like to thank our thesis advisor Dr David Saiang of the Department of Civil, Environmental and Natural Resources Engineering at Luleå University of Technology. We would also like to acknowledge Dr Oleg Nikitin at Enefit Kaevandused AS as the second reader of this thesis, and we are grateful for his very valuable comments on this thesis. Finally, authors of this thesis would like to acknowledge Enefit Kaevandused AS for their part in making this project come to fruition. 7

9 1. INTRODUCTION Oil shale is the most important energy resource of Estonia and related industries are biggest job providers in the country. There are three big energy companies involved in the oil shale mining and processing industry: Eesti Energia AS, Viru Keemia Grupp AS and Kiviõli Keemiatööstus OÜ. These companies employ about 9500 people and the whole sector contribute about 4% of the national cross domestic product [1] [2]. This Master s degree project is done in cooperation with the Enefit Kaevandused AS. 1.1 Background The bedding depth of the commercial oil shale bed is increasing southwards in the Estonia deposit. In Estonian practice, the economically reasonable depth of oil shale bed for opencast mining is up to 30 m. For greater depths than this, underground mining methods are employed. The mining front in Narva quarry is approaching this limit, i.e., the thickness of overburden is reaching 30 m. Also, in the SW part of the quarry the oil shale reserves are covered with peat reserves, and there underground mining might be the only way to extract the oil shale resource without destroying the upper peat layers. Enefit Kaevandused AS is planning to test underground mining (longwall mining) method as a potential technology for oil shale extraction for greater depths. Longwall mining technology has become significant in connection with the Enefit Kaevandused AS plan to deploy the technology (longwall mining) in the southern part of the Narva quarry mining field. In the first decade of the 21st century, Viru Keemia Grupp at the time considered to introduce longwall mining in the yet to be developed Ojamaa mine, but most likely due to the outbreak of the global economic crisis in 2008, it was decided to go with a safer version, with the already well-known room-and-pillar mining technology [3]. Longwall mining could be considered for implementation at the prospective Uus-Kiviõli oil shale mine. The choice for different mining methods in the Narva quarry southern part depends on the thickness of the overburden. With the existing draglines (EŠ15/90; EŠ10/70) the technological overburden boundary thickness is 27 meters, this means that with thicker overburden the mining technology should be modified. There are different solutions: it is 8

10 possible to introduce draglines with bigger parameters; increasing the thickness of the portable overburden with double-excavation method or use auxiliary machines. These kinds of changes will increase the production costs, leaving the mined quantity of oil shale the same, which means that no increase in productivity, only in the transferable overburden volume [4]. Longwall mining allows for extraction of oil shale from the ground without removing the overburden as such the production costs can be kept under control. Stable overlying rock thickness is important in the event of underground mining, in the case with room-and-pillar mining this should be more than 10 meters. In the southern part of Narva quarry however, such area is limited and it is only possible to use longwall mining [4]. 1.2 Project organisation This project is based on the data from previous studies and literary sources, geological data from Geological Survey of Estonia and data collected from numerical analysis carried out inside the framework of this study. The work here is carried out by 2 persons who are to analyse different problems within the same mining project: Taavi Lõhmuste stability analysis of the chain pillars using empirical methods, determination of optimal and safe chain pillar size, interpretation of the corresponding results; Ott Oisalu stability analysis of the highwall slope in punch-longwall mining using numerical methods, numerical modeling related to highwall stability, determination of optimal and safe barrier pillar dimensions, interpretation of the corresponding results. Geological and rock mechanical data is shared between the authors of this project and thus relevant investigation and work is shared between the parties. 9

11 1.3 Tasks of the project The main tasks of the project are: Study the geology of the site in order to determine the geotechnical parameters of the ore body and the overlying rock use drill core data, make appropriate maps of the area and geological profile of the ore and overlaying rock. Using MapInfo software for location maps and Bentley MicroStation for profiles; Determination of optimal and safe chain pillar size between highwall and panel end longwall face using empirical methods; Highwall stability calculation for Punch longwall using numerical methods; To determine stable and safe barrier pillar dimensions between the final longwall position and the highwall slope using numerical methods; Suggestions for highwall reinforcement methods. 10

12 2. METHODOLOGY 2.1 Baltic Basin oil shale resources Among the variety of big oil shale resources in the world the oil shale in the Baltic Basin is one of the most unique with respect to its composition and quality, which is among the top. The Baltic oil shale is also relatively easy to mine. That is why oil shale has been mined here almost the whole 20 th century Characterization of oil shale in Baltic basin The largest part of the Baltic oil shale basin is formed by Estonia and Tapa Deposits on Estonian territory (Figure 1). Tapa Deposit has so far not had any industrial importance. Estonia Deposit has been researched enough and extensive mining operations have been taken place. The oil shale in this deposit is called kukersite to stress its originality. It comes from the German name of the manor house in the neighbourhood Kuckers [5]. The oil shale seams occur among the limestone seams in Kukruse Regional Stage of the Middle Ordovician (the thickness of the seams extends from some centimeters to 0.9 meters). The thickness of the layer is 3-23 meters and the number of oil shale seams reaches 20 among which the lower ones form Eesti Deposit and the upper ones form Tapa Deposit. Territories of the deposits do not practically overlap with one another [5, 6]. Oil shale, generally speaking, is an organic-rich fine-grained sedimentary rock containing kerogen (a solid mixture of organic chemical compounds) from which liquid hydrocarbons called shale oil (not to be confused with tight oil - crude oil occurring naturally in shales) can be produced. Shale oil is a substitute for conventional crude oil; however, extracting shale oil from oil shale is costlier than the production of conventional crude oil, both financially and in terms of its environmental impact. Deposits of oil shale occur around the world, including major deposits in the United States. Estimates of global deposits range from 4.8 to 5 trillion barrels ( to m³) of oil in place [7, 8]. Oil shale, in Estonia, is a yellowish-brown, light, relatively soft sedimentary rock which contains a great amount of carbonate fossils. The main components are organic matter or kerogen (15-46%); carbonate matter (26-57%) and clastic material (18-42%). The main content 11

13 of the organic matter is formed by the microscopic round colonies of blue algae which are dispersed in the carbonate matter. Oil shale bed in Estonia deposit (production bed) contains six oil shale seams alternately with the limestone partings. The seams are marked with Latin letters A, B, C, D, E and F, beginning with the lowest seam (in more detail below). The oil shale seams located higher than F are not exploited [5, 6, 9]. Oil shale is easy to break and bore mechanically. The compressive strength of oil shale is MPa and that of limestone is MPa. The compactness of oil shale is t/m³ and that of limestone is t/m³. The characteristics of certain oil shale seams are quite different. The calorific value of the dry material is about MJ/kg ( kcal/kg) according to the seam and the area in the deposit. It is mainly influenced by the number of limestone inclusions (the amount of inclusions in the seams A-E is about 11%; in seam F, it is 39%) [9]. It is remarkable that in oil shale the relation between H atoms and C atoms is at least two times more than in the organic matter of coal. That is why the oil yield from oil shale by thermal processing is relatively high (65-70 % of organic matter). Kukersite contains a relatively little amount of sulphur which mainly occurs in a sulphide form [6]. Limestones under and above the production bed quarantees a strong and a safe floor and roof to the mine workings. The whole complex of Ordovician limestone and the Kukruse Regional Stage is penetrated by vertical tectonic fractures. They are mostly filled with argillaceous and sandy material. They allow water to easiliy penetrate through the rock massive, which leads to dissolution of soluble rocks as limestone and oil shale. Resulting landscape is called karst topography or just karst. These zones in Estonia deposit usually do not contain any oil shale, also their rock mechanical features are unpredictable and inferior to regular geologic composition. Areas like that need much greater support systems [5, 9]. The massive limestone is very damp as it continuously gets water from precipitation. It is necessary to pump out approximately 10 to 20 m³ of water per every oil shale tonne mined. This water contains fine particles from mining activity and due to law and environmental reasons needs to be treated before it is inserted back into the hydrosphere [5, 9]. 12

14 2.1.2 Kukersite distribution in Estonia and geological strata In the previous chapter it is stated that Estonian oil shale is named as kukersite or kukersite oil shale. Kukersite is a sedimentary rock that is part of a Kukruse stratigraphical stage of Late Ordovician period [9]. From economical and technical aspects only the lower parts of the Kukruse deposit are suitable for mining (see Appendix 1). Valuable layers are marked with letters A...F and limestone layers between them with combination of appropriate oil shale markings (underlaying/overlaying) E/F...A/B, or with names: E/F Pink limestone; C/D double limestone; A/B blue limestone. Some oil shale layers are divided into subcategories and marked accordingly. For example, for F layers the lower part is marked with index F1 and the upper part with F2. For the A layer, the upper part is marked with A. The production zones are combinations of A...F layers. They dip towards the north-east with the maximum depth located in North-Eastern part of the country. As previously mentioned the production zones form the Estonia deposit (Figure 1). A...F layers and overlaying G- and H layers are thickest in Jõhvi hill chain, located South-East from city of Kohtla-Järve toward middle of the deposit. The thickest parts of oil shale also contain the highest amount of organic component called kerogen. Overlaying the production zone are G...P layers which are thin and separated from each other with relatively thick limestone layers. Above P are layers which are marked with Roman numerals I...VII and they form the Tapa deposit (Figure 1). Oil shale found on Tapa deposit is economically not minable today (low organic compound content and at great depth) and thus is also not suitable for longwall mining method [9]. 13

15 Figure 1. Oil shale distribution map [10] Characterization of Estonia deposit Alternating oil shale and limestone layers are distinguishable by thickness, strength, composition, hardness, density and heat of combustion (see Table 1, Appendix 4). Larger kerogen content inside the layer ensures higher heat of combustion (see Table 1). In addition, the thickness of the layer is also very important. The B layer by a rule is the thickest and contains the largest amount of kerogen, which means it has a large heat of combustion value. The thickness of the layer is about 0.80 meters in the middle of the deposit and decreases towards periphery [6]. In the middle of the production layer lies the C/D limestone layer with a thickness of about 0.25 meters. This is the strongest limestone interlayer with uniaxial compressive strength slightly over 80 MPa (see Appendix 4). The E layer has the highest kerogen content after B and has a thickness of about 0.4 meters [6]. 14

16 Table 1. Average quality characteristics of oil shale in Estonia deposit [9]. Layer Heat of combustion Q Kerogen content K kcal/kg MJ/kg participation % F F E D C B A' A The thickness of Estonian deposit production layers vary between 2.5 and 3 meters, where the oil shale part is meters and limestone around meters [6]. The Kukersite layers are located near the surface in the Northern side of the deposit and these favor the strip mining method. In order to better illustrate the oil shale production layers in an Estonian deposit a histogram is presented (see Figure 2, Appendix 2), which shows geological composition of oil shale by sub-layer thickness among different production and research fields. From Figure 2 one can determine that the production layer is relatively similar along the whole deposit, the average thickness is around 2.7 meters. Crystalline basement and sedimentary cover dip towards the south, about 3.5 meters per kilometer with the production layers also dipping accordingly [6]. This means that as the recoverable oil shale goes deeper, underground mining methods will be more feasible. This thesis focuses on the Narva mining field where Enefit Kaevandused is planning to use longwall mining technology. Further geological study concentrates primarily on that region. 15

17 Thickness [m] Layer index Production layer thickness in Estonia deposit 3 F E D/E D C/D C B/C B A1/ B 0 Estonia Narva Sirgala Ojamaa Uus-Kiviõli Mining field/research field Figure 2. Production layer thickness in Estonia deposit [11] Narva mining field backround information and location description Enefit Kaevandused AS owns four mining licences on Narva and Sirgala ore fields: Sirgala open cast KMIN-074, Narva open cast KMIN-073, Narva II oil shale open cast KMIN- 046 and Sirgala II oil shale open cast KMIN-087. Work along previously named mining shares is carried out uniformly as one production area and for simplification is named generally as Narva open cast (see Table 2) [12]. Enefit Kaevandused AS has submitted an application to the Estonian Ministry of Environment in order to unify the annualy allowed maximum mining rates in above-mentioned licences to one uniformal licence. [13] The maximum annual allowable mining rate in Narva open cast is 6.4 million tonnes of oil shale (see Table 2). According to the balance of mineral resources the mineable reserve by end of 2015 was million tonnes. 16

18 Table 2. Narva open cast mine mining share [12]. Licence nr Mining share name Annual allowed maximum mining rate, thousand tonnes Licence validity Mineable reserve, thousand tonnes (2015) KMIN-046 Narva oil shale open cast II KMIN-073 Narva open cast KMIN-074 Sirgala open cast KMIN-087 Sirgala II oil shale open cast Narva open cast mining shares lie on eastern part of Estonian deposit, west of Narva River. Mining fields can be seen in Figure 3. [14] They are in Ida-Viru County, on Vaivara- and Illuka Parish lands. Figure 3. Narva open cast mine location and bore holes used in this thesis. 17

19 Narva oil shale quarry geology The Narva quarry is characterized by a relatively simple geological structure. Geological section, opened by mining, consists of (top to bottom) Quaternary sediments, middle Devonian Narva regional stage and upper Ordovician Kukruse regional Stage rocks. Narva regional Stage consists of clayey dolomites and marl - intermediate layers of clay and sandstone, with the average thickness of 14.3 m. Kukruse regional Stage sedimentary rocks consist of inconsistent clayey limestone layers with thin intermediate layers of marl and shale, at the bottom of the regional stage is the oil shale productive bed. Kukruse Regional Stage has an average thickness of 9.7 m. Quaternary sediments are represented with the subsequent ice age glacial and limnoglacial origin sandy-clay-pebble sediments, as well as Holocene lakemarsh and wind-borne sediments. Lake-marsh sediments pose as a semi-decomposed peat moss and everything is related to Puhatu bog massive. The layers dip slightly to the South. Significant tectonic faults do not occur in the mining field [15]. The compactness of oil shale is t/m³ and that of limestone is t/m³, as mentioned previously [9]. The oil shale industy productive bed further includes in addition to oil shale layers, also limestone layers and inclusions, which has a compressive strength of MPa, while that the oil shale is MPa [16]. The productive bed in Narva quarry consist of the Kukruse Stage lower part of seven kukersite (from bottom-up A, A, B, C, D, E, F1), and six limestone layers (A/A1, B/A, C/B, D/C, E/D, F1/E). Limestone layers A/A1 and F1/E are, in some spots, missing. The thickness of the productive bed is up to 3 meters. Kukersite layers thickness is different - from 0.05 to 0.6 meters. The thickness of limestone layers is up to 0.3 meters, and their contacts with oil shale layers are relatively uniform. The overall thickness of the productive bed, regarding the surface area of it, is well-kept [17]. In this thesis drill holes no (KMIN-073) and no (KMIN-046) were used to make the geological cross-sections (Figure 3). The results obtained are given in the following outlined schemes (see Appendix 3; Figure 4, Figure 5): 18

20 Figure 4. Thickness of production layer between boreholes PA 4260 and PA 4369 on section I- I from northwest to southeast (Figure 3). Figure 5. Production layer cross-section from northwest to southeast on section I I (Figure 3). 19

21 For both boreholes, the oil shale and limestone interlayers are quite similar, the contacts between the layers are straight (Figure 4). The productive bed dips towards the south about 2.78 meters. This can be seen in Figure 5 as the oil shale layer A' absolute depth difference between the boreholes PA 4260 and PA 4369 is 2.78 m. The thickness of the productive bed is 2.68 meters and 2.54 meters; limestone layers thickness is 0.74 meters and 0.72 meters respectively (see Appendix 3). On top of borehole PA 4260 productive bed lies 25 meters of overburden, meters in case of borehole PA 4360 (see Table 3). Stripping ratio increases with productive bed depth increase (see Figure 6). Figure 6. Thickness of overburden layers and structure on section I-I (see Figure 3). 20

22 Table 3. On top of the productive bed lying overburden parameters [18]. Overburden Borehole no. Borehole no. Description Description Overburden total thickness (Q+D2+O2) [m] Q1 [m] 0.8 Peat 0.4 Peat Q2 [m] 12 Sand with pebbles and cobbles 7.1 Sand D2 [m] 5 Marl, dolomite 12.3 Fissured marl, dolomite O2_F2 [m] 7.2 Dolomitic limestone 7.94 Limestone 2.3 Geotechnical model This section provides information about rock mechanical parameters used in this thesis and builds the so called geotechnical model to be used in modelling later in FLAC and RS 2. Knowing the mechanical properties of the rock masses above the extracted oil shale seam is very important to predict the accompanying deformations from roof caving, stress distributions and loads to the support system. The mechanical properties are in turn controlled by the strength properties of intact rock, but also by the presence of discontinuities in the rock mass as these form weakness planes. [19] The footwall, commercial oil shale bed and overburden were divided into different layers in the model. The model was built up by knowing two simple variables: the thickness, layout of different layers and the geotechnical properties of each layer. The thickness data for the model was aquired from the previous chapter (Ch Narva oil shale quarry geology). 21

23 Also, some more precise geological information was given by the Enefit Kaevandused AS. The rocks and soil properties, as well as other model parameters were given by Enefit Kaevandused AS. The properties were aquired from previous studies, observations of roof collapsing tests, old mine roof behaviour and from the current outcrop field work. The mining company ordered laboratory tests to confirm the reliability of the most important properties (e.g. compressive strength). Hoek & Brown failure criterion was applied in the model. Rock masses which have a sufficient number of closely spaced discontinuities and similar surface characteristics to assume the isotropic behaviour involving the failure of the discontinuities. The rock mass can be treated as Hoek&Brown material, when the structure that is being analysed is large and the block is small in comparison. [20] Description of soils and rocks The geotechnical model contains six layers: overburden (soil and rock overburden), main roof, immediate roof, commercial oil shale bed and footwall. This layout can be seen in Table 4. The overburden covers the commercial oil shale bed. The upper part of the overburden is composed of Quaternary soils and the lower part of sedimentary rocks. The overburden consists of Quaternary cover and bedrock. The Quaternary cover comprises of Layers 1-5: fill layer, topsoil and peat, sand, clayey silt/silty clay, till and other clayey sediments. In the geotechnical model the Quaternary soils are merged into single unit (soil overburden), because the differences in soil properties are insignificant. Quaternary soils properties are important when considering the effects on hydrogeological conditions. The thicknesses of Quaternary cover and the bedrock overburden can be seen in previous chapters. The roof in Estonian oil shale mining practice has been observed to cave in two stages. As the caving is gradual, the roof is divided into two layers: immediate and main roof. The distinction between these two units is based on deformation and caving behaviour during mining. In here, the first caving layer is referred to as the immediate roof and the material that caves after that is referred to as the main roof. The previous experience has shown that the immediate roof approximately coincides with the lowermost part of bedrock overburden that consist of altering layers of limestone and non-commercial kukersite oil shale (Layer 8). The main roof usually comprises the remaining upper part of Kukruse Stage limestones (Layer 7). 22

24 Bedrock overburden Table 4. Scheme of geotechnical units in the model, their geological classification, type and layer number. Geological unit in model Geological unit Soil and rock type Layer number fill and topsoil Layer 1 peat Layer 2 Soil overburden Quaternary cover sand Layer 3 clayey silt, silty clay Layer 4 till Layer 5 Narva Stage Narva Stage marl and dolomite Layer 6 Main roof limestone Layer 7 Kukruse Stage limestone with Immediate roof kukersite Layer 8 Mined oil shale bed kukersite oil shale Layer 9 Footwall Uhaku Stage clayey limestone Layer Description of discontinuities The most common strikes of major sets of joints in Ordovician carbonates are 320 and 53, the joints are open with the width of aperture in the range of few mm to few cm [21]. More specific studies in NE Estonia [22] also point out similar trends: the zones trending NW and NNE 5-25 usually form narrow river valleys or linaments; the NNE-trending fault zones occur mostly in the eastern part of the oil shale deposit (e.g. in Narva quarry area). Based on the distance between joints the rock masses can be divided into two zones: normal zones and highly fractured zones. The normal zones are most widely spread. Adjacent joints spacing averages m, but can also be greater. Most of the joints are not filled. The rock face is monolithic and undisturbed. Normally these comprise about 90% of the observed outcrop. [23] The highly fractured zones are the areas where the joint spacing has decreased to m or even less than 0.1 m. The distance between joints usually decreases gradually towards the center of the zone. The structure of rocks can be slightly disturbed, especially in the center of the zone. Also joints may exhibit more or less clayey filling. The average width of the highly fractured zones was about m and for every m of observed outcrop at least one fractured zone was present. Of course the actual distribution of different zones does not form that regular pattern. [23]. 23

25 2.3.3 General geomechanical properties of rock The Hoek-Brown criterion requires mechanical properties as an input including unit weight, properties of intact rock and rock mass rating. The latter combines the intact rock parameters and properties of discontinuities in order to estimate the strength and stiffness of the rock masses. Table 5 lists the general geomechanical properties of the intact rocks. These properties are used in the model to calculate stress conditions and specific Hoek-Brown parameters. The presented values are combination of literature study and laboratory testing conducted in 2013, ordered by the Enefit Kaevandused AS. In this thesis summary of these values were given by the company. The key parameters in Hoek-Brown criterion are uniaxial compressive strength and modulus of elasticity, thus they were chosen to be tested in laboratory in order to evaluate the reliability of values provided in the literature. The most important layers in the context of modelling are the roof strata, thus the rocks from overburden were tested. The comparison of laboratory results and literature is shown in Table 6. The laboratory testing of elastic modulus resulted in unrealistically low values which do not characterize the intact rock, but disturbed and fractured mass and thus cannot be used directly in the model. The values from literature were used instead, because the Hoek-Brown criterion requires the elastic modulus of intact rock. As suggested by Hoek [20], in general, the reliability of measured value of elastic modulus is suspect because of specimen damage. The specimen damage has a greater impact on modulus than on strength and hence, the intact rock strength should be used to calculate elastic modulus [20]. The elastic modulus was calculated using equation [20]: E = MR σ ci (1) MR modulus ratio; σ ci uniaxial compressive strength of intact rock Modulus ratio values were chosen according to scheme in Figure 7. 24

26 Layer number Table 5. General geomechanical properties of 6-10 layers. Geotechnical unit Rock type Unit weight γ, kn/m³ a General properties Uniaxial Compressive Strength σ ci, Mpa a Poisson' s ratio ν c Modulus of Elasticity E, Mpa b Modulus ratio (MR) marl 23,5 42 0, Narva Stage dolomite , Main roof limestone , Immediate roof 9 Commercial oil shale bed limestone with kukersite kukersite , , clayey , Footwall limestone a values from laboratory testing/from references [24] b calculated values [20] c values from [25] For modelling purposes the Quaternary sediments (Layers 1-5) are considered as one geotechnical unit with Mohr-Coloumb failure criterion (unit weight γ = 22 kn/m³, friction angle φ = 33, cohesion c = 5 kpa) and Young s modulus E = 20 MPa. The elastic properties and type of the quaternary sediments material is defined as isotropic and elastic. 25

27 Layer number Table 6. Comparison of the values of key parameters from references and laboratory testing. Unit weight γ, kn/m³ Uniaxial Compressive Strength σ ci, Mpa Modulus of Elasticity E, Mpa Geotechnical unit References Laboratory References Laboratory References Laboratory 6 Narva Stage, marl Narva Stage, dolomite 23,5 a 22,5-23,5 49 a a ,5-24, Main roof 23,5 a 24,5-25,5 69 a a Immediate roof 23,5 a 24,5-25,5 20 (oil shale) (limestone) a values from [24] (oil 33 a shale) (limestone) Figure 7. Values for rock mass modulus [20]. 26

28 2.3.4 Geomechanical properties of rock Hoek & Brown parameters After determining the general geomechanical properties, the specific Hoek-Brown parameters can be calculated. These parameters depend on the properties of intact rock, but also take into consideration the pattern and characteristics of discontinuities present in the rock mass [20]. The above-mentioned properties of rock mass for normal and for highly fractured rock mass, are listed in Table 7 and Table 8. The concept and evaluation of the Hoek-Brown parameters is described subsequently. Rock parameter (mi) depends on the texture and composition of a rock type. According to the scheme [20] in Figure 8 the rocks in Narva quarry are mainly different types of carbonates, but the oil shale layer was evaluated to be between claystone and shale, as Estonian kukersite oil shale is not actually a traditional shale. The structure (from intact to laminated/sheared) of a rock is determined by the discontinuities pattern in the rock [20]. In normal monolithic zones the structure can be described mainly as blocky, the limestone layer is occasionally intact. Geological Strength Index (GSI) is determined by the structure and discontinuities surface characteristics, e.g. whether the surfaces are smooth or rough [20]. The higher the GSI value, the stronger the rock mass. This concept is presented in Figure 9. The Hoek-Brown constant (mb) and rock mass characteristic constants (s and a) which describe the rock mass strength, were calculated using following formuli [20]: (2) if GSI > 25, then s and a are: (3) a =

29 Figure 8. Values of the rock parameter (mi) for intact rock masses [20]. Figure 9. GSI (Geological Strength Index) estimation based on the geological descriptions [20]. 28

30 Layer number Layer number Table 7. Hoek-Brown parameters of normal monolithic rock. Geotechnical unit Rock type Rock group m i Hoek-Brown parameters of monolithic zones Uniaxial Hoek- Strength Compr. Brown Structure index Strength constant GSI σ ci, Mpa Rock mass characteristic constants m b s a marl 6 Blocky ,01 0,0039 0,5 6 Narva Stage dolomite 9 Blocky ,80 0,0067 0,5 7 Main roof limestone 10 Intact/blocky ,43 0,0357 0,5 8 9 Immediate roof 10 Footwall limestone with kukersite Commercial oil shale bed kukersite clayey limestone 7 Blocky ,17 0,0039 0,5 5 Blocky ,84 0,0039 0,5 9 Blocky ,51 0,0039 0,5 Table 8. Hoek-Brown parameters of weaker highly fractured zones. Geotechnical unit Rock type 6 Narva Stage marl dolomite 7 Main roof limestone limestone 8 Immediate roof with kukersite 9 Commercial oil shale bed 10 Footwall kukersite clayey limestone Rock group m i Hoek-Brown parameters of weak zones Uniaxial Hoek- Strength Compr. Brown index Strength constant GSI σ ci, Mpa m b Structure very blocky very blocky very blocky very blocky very blocky very blocky Rock mass characteristic constants s a ,70 0,0013 0, ,06 0,0013 0, ,17 0,0013 0, ,82 0,0013 0, ,59 0,0013 0, ,82 0,0013 0, Model parameters rock mass properties The rock mass properties are estimated using the RocData software (Rocscience Inc., Canada), based on the geomechanical properties of intact rock pieces tested in the laboratory and rock mass characterization (see Chapters and 2.3.4). 29

31 Input parameters to RocData include uniaxial compressive strength of intact rock, geological strength index, intact rock parameters, disturbance factor and Young s modulus of intact rock. Due to blasting in the Narva open cast mine, the highwall stability is influenced by that, however the blast damage is usually restricted to the first few meters of rock. So, the damage zone material will be incorporated into numerical model as a different and weaker material than the surrounding rock mass. Due to non-blasting extraction, in the case of longwall mining, the disturbance factor in this case is selected to be zero. The elastic properties and type are defined as isotropic and plastic. The results of the RocData output, i.e. rock mass parameters used in modelling are presented in Table 9 and Table 10. Table 9. Rock mass properties for modelling D = 0 (normal rock mass). Layer Cohesion c, Mpa Internal friction angle φ, Tensile strength σ t, Mpa Uniaxial Compressive strength σ c, Mpa Global strength σ cm, Mpa Deformation modulus of rock mass E m, Mpa 6 - marl 1,8 26,3-0,16 2,53 5, dolomite 3,4 31-0,24 5,23 12, ,9 36,3-0,72 12,97 19, ,5 27,5-0,11 1,99 4, ,3 45-0,11 1,5 3, ,6 29,6-0,14 3,31 9, Table 10. Rock mass properties for modelling, highly fractured zones (used for DZ - damage zone). Layer Cohesion c, Mpa Internal friction angle φ, Tensile strength σ t, Mpa Uniaxial Compressive strength σ c, Mpa Global strength σ cm, Mpa Deformation modulus of rock mass E m, Mpa 6 - marl 1,5 23,4-0,07 1,38 4, dolomite 2,6 26,7-0,07 2,15 8, ,9 27,5-0,07 2,28 9, ,2 24,7-0,05 1,09 3, ,8 22,1-0,05 0,82 2, ,3 51-0,08 1,82 6,

32 2.3.6 General model description In order to perform numerical modelling, the mechanical properties of the immediate roof (Kukruse Stage limestone and kukersite layer 8), floor (Uhaku Stage clayey limestone layer 10), commercial oil shale seam (Kukruse Stage layer 9), roof (Kukruse Stage limestone layer 7) and other overlaying rocks (Narva Stage dolomite and marl layer 6) have been determined. The Hoek-Brown rock failure criterion has been selected for the modelling process. In Table 11 the units and layers of the model are re-introduced with some remarks. Rocks (Layers 6-10) are considered to have isotropic elastic properties and they represent plastic type of material. The state of in-situ stresses is considered to be σv = γ * Z and σh = 0,33 * σv in vertical and horizontal planes, respectively. This horizontal-to-vertical stress ratio of rocks is lower than proposed by common theory [26], but the lower value appeared to match better with the observed behavior in the mining practice [24]. According to the common theory [26] the ratio of horizontal to vertical stress increases with the decreasing depth. Although, the near surface horizontal stresses are highly influenced by effects of nearby topography or a high degree of near surface fracturing [27]. In Narva quarry the mining depth is too shallow for the parameters suggested by Sheory [26]. In the study area, this could be explained by the impact of postglacial rebound that has occurred during the last years after the last glaciation event. The postglacial rebound causes the release of stresses and thus the horizontal-to-vertical stress ratio can be lower. Groundwater level is supposing to be below the mining level due to drainage. 31

33 Table 11. Geotechnical units in the model with some remarks. Geological unit in the model Quaternary overburden Narva Stage Layer number Layers 1-5 Layer 6 Description and remarks Quaternary soils form a single unit in the model, because the difference in soil' properties are insignificant. All the soils have isotropic elastic properties and are considered as elastic type of material. Rather thin-bedded; layers of marl, dolomite and even clay are alternating. Not rational to include all the small layers in the model. Using smaller number of layers. The upper 1/3 of Narva Stage is defined as marl in the model, the other 2/3 is defined as dolomite. In order to better imitate the layered nature of Narva Stage, thin marl layer (thickness 0,5 m) was inserted at the bottom of Narva Stage. Main roof Layer 7 Composed of Kukruse Stage limestone. In order to better describe the layered nature, a thin (0,5 m) layer of marl was inserted dividing the unit into two blocks. Immediate roof Layer 8 Although the immediate roof is divided into thin oil shale and limestone sub-layers in reality, for model simplification it was considered as a single layer. The weighed average compressive strength of these layers was used in the model. Mined oil Layer 9 Commercial kukersite oil shale of Kukruse Stage shale bed Footwall Layer 10 Clayey limestone of Uhaku Stage Figure 10 shows the modelled highwall slope geological cross-section with the elevation above sea level data. On the figure each layer thickness is marked with a white box. This was used as a base for creating models in RS 2 and FLAC. The quaternary cover is not in the model, because it is removed in the stripping stage by draglines to expose the bedrock. The thickness data was acquired from Table 3 as an average of those boreholes data. The exact place of the cross-section (marked with a blue line) can be seen in Figure 11. Ground surface caving is expected when using the longwall mining method. In general, the overburden above mined seam is divided into three zones based on their deformations: caved zone, fractured zone and continuous zone [28]. In the caved zone, the rocks fall to the mine floor and are broken into irregular pieces; according to different data from the literature, thickness of this zone varies between 3-12 times the thickness of mined seam [28]. Above the caved zone is the fractured zone that is characterized by fractures caused by mining and its 32

34 thickness varies greatly between 20 to 100 times the mined seam thickness [28]. In the continuous zones no fracture or other significant deformations occur. In Narva quarry the depth of oil shale is so shallow that the whole overburden forms caved and fractured deformation zones. The highwall stability design was modelled using a 10-stage FEM model. Stage 1 characterizes the conditions, where 7 m long chamber for shearing machinery, conveyor and support system is prepared. After that the longwall advances with 5 m steps. Various stages can be seen in Figure 13 and Figure m Marl 3.5 m Dolomite 4.5 m 2.5 m 3.5 m 3 m 2 m Figure 10. Cross-section of the highwall [modified according to 14] (see Figure 11). 0.5 m marl layer 33

35 Figure 11. Narva open cast mine (blue line indicating a geological cross-section location and green line indicating a hydrogeological cross-section location) Hydrogeological description When oil shale is mined, the above lying aquifer is completely drained in the mining front. The bedrock has a southward inclination and when the mining front moves southwards, the drawdown increases and so does the impact of the drawdown. The drawdown in Keila- Kukruse aquifer is the main factor impacting the adjacent areas. It forces the groundwater to flow into the depression and lowers the water table in the areas outside of the borders of mine. There is no difference, if the mining is performed in open pit or underground, because the drawdown in both cases is the same. The upper part of the hydrogeological section (two uppermost aquifers) was included in the analysis, because these upper aquifers are mostly affected. The upper part of the hydrogeological section is divided into two aquifers: unconfined aquifer in Quaternary sediments (peat, sand) 34

36 confined aquifer in limestone and dolomites Keila-Kukruse aquifer. The aquifers are separated by the confining unit consisting of the Quaternary clayey soils (clayey silt, silty clay, glacial till) and Middle-Devonian complex of marls and dolomites of the Narva age. The Quaternary aquifer is formed in sands of glaciofluvial and glaciolacustrine genesis. The overlying peat is also included into the Quaternary aquifer. The Quaternary sand and gravel form an important aquifer west of the mining area (Figure 11). In the Vasavere ancient valley within the Kurtna kame field sand and gravel form an up to 80 m thick layer. The aquifer recharges directly from precipitation, discharge occurs into the open pit. [29] Moving to the east to the study area (Figure 11) the aquifer thickness decreases, remaining between 3-6 meters. In the mining area the glaciolacustrine sediments fine sands with some silt are dominating, locally crossed by zones of the fluvial sand and gravel. [29] Keila-Kukruse confined aquifer comprises of limestones and dolomites, the commercial oil shale is also counted into the aquifer. The thickness of the aquifer, in the distribution area of Narva stage, is about 15 m, the thickness increases moving to the west (Figure 11), reaching up to meters. The aquifer is missing in the Vasavere ancient valley which is cut deep into the rocks of the Uhaku stage. The recharge area for the aquifer is located in the Kurtna kame field and Vasavere ancient valley where the groundwater table lies at altitudes m (a.s.l). It creates the pressure in the aquifer and where the confining unit is present, the water level of the aquifer rises above the top of the aquifer. [29] The cross-section shown in Figure 12 illustrates the hydrogeological situation in the region of trenches 8-12 (see thick green line in Figure 11), where the mining front locates farthest in the south and the drawdown in the Keila-Kukruse aquifer is maximal. The groundwater level of the Keila-Kukruse aquifer in the time being is shown by light blue dashed line (Figure 12). Quaternary aquifer level is shown by dark blue dashed line. Figure 12 was used as a template for the groundwater level data for numerical modelling. 35

37 Figure 12. Hydrogeological section [modified from 6, 9, 18, 21] Model description for the highwall RS 2 In order to study the stability of the highwall slope a model was constructed. The model was simulated using the elasto-plastic finite element stress analysis program RS 2. Plastic analysis is used in order to get information about yielded elements and to get a glimpse of the behaviour of the slope. Linearly elastic-perfectly plastic material properties were used in the plastic model. Groundwater level was implemented to the model by using data from previous chapter. The model external boundaries were constructed, keeping in mind that model boundaries should be sufficiently large enough so that disturbances to the stresses due to mining are within the model boundaries. The extent of the modelled domain was defined by an expansion factor of 4-13 in relation to the slope height, to eliminate boundary effects. The domain was discretized with a uniform mesh of six-noded, triangular elements. Approximate number of mesh elements was picked to be 5000, as can be seen in Table 12. In the model, gravitational field stress was used (Table 13) as in-situ stresses are deduced by an educated guess. 36

38 Table 12. RS 2 input parameters description. Properties Model Limitations Boundary Custom boundaries: 13 times the height of the slope in sides; vertically 4x times the height of the slope Displacement Top is free, fixed on bottom corners; roller support on sides and bottom Stress field Gravity Initial element loading for undisturbed Field stress and Body force rock mass Mesh Mesh type Uniform Element type 6-Noded Triangle Gradiation factor 0.1 Number of nodes on all excavations 5000 (Mesh density increased 4x times around the slope area - the discretization and mesh element density) Table 13. RS 2 stress state input data. Input data Stress data Vertical stress σv=ρgz Total stress ratio (in-plane) 0.33 Total stress ratio (out of plane)

39 Figure 13. Highwall slope model in RS 2. Blue line indicates a groundwater level. The model described above is shown in Figure 13. The slope angle was set around 80 degrees, as it is close to the actual slope angle in the open pit. As can be seen the mesh around the slope is multiplied 4 times the usual mesh density. Also, the SSR Search Area is applied in the model. A Shear Strength Reduction (SSR) analysis is carried out with RS 2, which by default considers the stability of the entire model when the analysis is computed. The search area was applied to the highwall slope region. SSR itself is a finite element slope stability analysis and computes a critical strength reduction factor in the model. The critical strength reduction factor is equivalent to the safety factor of the slope. This can be compared to the FLAC model factor of safety value Model description for the highwall FLAC For assessing the overall highwall slope stability explicit finite difference numerical program (FLAC) was used beside RS 2. FLAC will help one determine the failure mechanism and identify the failure plane directly by reducing the strength of the material until failure occurs. Model in FLAC was created using the slope builder tool. There one can insert the slope parameters (height of the slope, angle of the slope, model depth, model width to right and left). The model depth and width was chosen to ensure that any failure surface, which develops 38

40 will not extend to a model boundary. The model depth was about 4 times the slope height and the width was about 15 times the slope height. Slope angle was chosen around 80 degrees. Boundary conditions was chosen to be roller boundaries on the sides of the model (fixed in the x-direction) and on the base (fixed on the y-direction). The upper part of the model was left without boundaries as it is a free surface. Model geometry and boundary conditions can be seen in Figure 14. The model was divided into a mesh of quadrilateral zones (Figure 14). Very fine zoning (300) was used to get a more accurate calculation. It provided a fast solution and was reasonably accurate. A Hoek-Brown constitutive material model was used and the following material properties were assigned for the model (Table 5; Table 7; Table 8; Table 9; Table 10). Groundwater level, that is used in the model, was based on Figure 12. The stress field in the model was assumed to have larger vertical stresses than horizontal, because postglacial rebound causes the release of stresses and thus the horizontal-to-vertical stress ratio can be lower. The vertical stress was assumed to be gravitational, so it was calculated using a simple relation of the unit weight of the overlying rock and the depth below the surface. Horizontal-to-vertical stress ratio of 0.33 was used to calculate the horizontal stresses. The stress state, which was used in all the models, was calculated as described in the previous paragraph. A weighted average density of the overlying rock was used, because of the layered rock type, for the modelling: ρw.ave. = 2366 kg/m³. Used in-situ stresses were following: σv = 1.85 Mpa and σh = 0.61 Mpa. The stresses were calculated in the bottom of the model. The out-of-plane stress was assumed to have the same magnitude as the horizontal stress. 39

41 Figure 14. Highwall slope model in FLAC. 2.4 Technology of (punch) longwall mining Longwall mining in Narva open cast is a mining technology that Enefit Kaevandused is planning to implement. Figure 15 and Figure 16 present a general plan view to exemplify how this method is going to be used. Shearer is cutting oil shale at the face and it rides on the panline of an armored face conveyor (AFC), which is located on the floor perpenticular to the mining 40

42 direction. The AFC provides track for the shearer to move on. The shearer is hauled by two self-contained electric motors on each end of the machine turning a series of sprocket wheels that run on specialized track set on the collapsed roof (gob) side of the chain conveyor. There are two rotating drums mounted on each end of the shearer that do the work and cut oil shale. The position of these drums can be hydraulically adjusted by the ranging arms. [30] Figure 15. Plan view of (punch) longwall mining method (two-entry system). 41

43 Figure 16. Plan view of (punch) longwall mining method (three-entry system). Broken rock consisting of oil shale and limestone is loaded onto the AFC and hauled to the headgate T-junction. Headgate is also a headentry or adit in this case, as in punch-longwall mining the entry to the mineable area is built from highwall straight into the production zone. A T-Junction is the intersection of the longwall face and the headgate or tailgate (tailgate is located on the side of the mined-out panel). Broken rock is then dumped from the AFC to the crusher and from there to the stageloader which empties it to the belt conveyor and lastly the crushed rock is taken out of the mine. [30] Shield supports are used to support the roof at the face so that equipment and work force are safe. They are set separately so that when the shearer cuts and passes then appropriate shield supports can move and immediately support the roof on top of the chain conveyor. As seen from Figure 17, both AFC and shield supports can move like a snake in order to follow the shearer. Shield supports will hydraulically push the conveyor forward. [30] 42

44 Figure 17. Movement of AFC and shield supports [30]. During panel developement, the panel entries/gateroads or adits, in this case, are roof bolted like in room and pillar mining [30]. The gateroads as well as the whole mine is supported by the chain and barrier pillars. Determining the size of these pillars is a crucial part of this thesis. Barrier pillars are also important for the stability of the highwall. Two different gateroad development systems are compared in this paper, two-entry system and three-entry system (Figure 15 and Figure 16). For two-entry system the first entry is used for the conveyor belt (and track or diesel car road) while the 2nd entry is the tailgate for the next or future panel [30]. For three-entry system the first entry is for the conveyor belt, the second entry is the track for travel, and the third entry is the tailgate for the next panel. [30] The advantage of longwall mining ahead of conventional room and pillar mining is higher productivity and smaller oil shale losses. On the downside there are higher initial capital costs due to relatively high price of the equipment needed. Also, there is a significant impact on the ground directly above the mine as the roof collapse causes surface subsidence. Figure 18 exemplifies the changes on the ground due to longwall mining. Subsidence in Estonia is around 1.8 meters as the thickness of the production layer is approximately 3 meters. [31] 43

45 Figure 18. Cross-section of longwall mining. 2.5 Chain pillar design Chain pillars are an important part of the longwall mining process as they influence the stability of the mine and entries, also the production levels are affected by the size of these structures. In this thesis mainly the stability aspect of chain pillars in a longwall mine is studied, also appropriate dimensions regarding safety are proposed. Analysis is done for Enefit Kaevandused AS in accordance with Narva mining field geological conditions. Empirical methods are most widely used for that purpose as they are based on the real life cases and can be easily adjusted to different geological and rock mechanical situations. Chain pillars offer stability and as the mine goes deeper or geological conditions are not favorable, then usually pillar parameters are increased in order to enhance the stability of the mine. This causes the overall oil shale recovery to decrease, thus mining costs per tonne 44

46 increase. Wide pillars hinder the air flow which also raise costs. According to this it is important to design the pillars as small as possible while maintaining adequate safety factor [32]. Optimum size of chain pillars in the Narva oil shale mine is designed in this thesis by using the analysis of longwall pillar stability (ALPS) empirical method Empirical method Empirical methods, based on case histories of similar mining techniques, are applicable in situations where good geotechnical input is available. Empirical methods are based on the basis of real life data and field observations, making them quite reliable when dealing with complex rock mechanical problems. Based on the necessity to make underground coal mining in United States safer and more dependable, a new method for the design of longwall chain pillars, Analysis of Longwall Pillar Stability (ALPS), was developed by C. Mark and Z.T. Bieniwski. ALPS produces approximations of the relative stability of different longwall pillar designs during all stages of their service life. This method consists of three main steps [33]: 1. Estimating the loading that will be applied to the pillars during all the phases of longwall mining 2. Estimating the load-bearing capacity of the longwall pillar system 3. Calculating Stability factors (SF) by comparing the load to the load bearing capacity. ALPS is an extensively used method and has been verified by back-analysis of more than 100 actual mining case histories [34]. In this thesis project it is used in accordance with Narva oil shale field geology and geotechnical data in order to get preliminary estimates of chain pillar parameters for longwall mining in Narva open-cast mine. One drawback of this method is that it does not take into consideration the effect of horizontal stress on pillar stability. Previous limitation is not important in this work because due to the shallow depth of oil shale in Narva mine the horizontal stresses are relatively marginal. Stability analysis of chain pillars in Narva oil shale mine is performed by using the ALPS software that was created by Mark, C. based on the ALPS method [34]. Mentioned computer program needs input data of longwall design parameters, loading condition and also user provided safety factor for making appropriate calculations and suggestions. 45

47 Input parameters for ALPS software are presented in Table 14. Panel specifications are taken from chapter and data acquired from Enefit Kaevandused AS (panel width 300 m, entry width 5.5 meters), chain pillar parameters are chosen by author for further analysis. Data for loading conditions is taken from chapters and Abutment angle is ~21 in United States coal mining practices and is used as a constant when using ALPS [35]. According to Enefit Kaevandused this should be similar when implementing longwall mining in Estonia. Further test mining needs to be done in order to determine it specifically. Table 14. Input parameters for ALPS Panel specifications Loading conditions Entry Height (m) 2.8 In-situ Shale Strength 3.16 Depth of Cover (m) 28 (Mpa) Panel Width (m) 300 Abutment Angle ( ) 21 Entry Width (m) 5.5 Chain pillar width (m) 4 to 8 Overburden Unit Weight Chain pillar length (m) 4 to 8 (kn/m 3 24 ) Number of Entries 3 ALPS software is used for determining optimal chain pillar parameters for longwall mining in Narva field by running the calculations with different chain pillar width and length values while other parameters remain constant. Chain pillar width and length of 4-8 meters are analyzed in this thesis paper. It is estimated by the authors and Enefit Kaevandused AS (based on experience from room and pillar mining) that pillar width under 4 meters may be unsafe and over 8 meters may be waste of resources. Analysis of this thesis aims to confirm this estimate and determine optimal pillar size. If it is found that other pillar parameters need to be analyzed, then corresponding analysis is added. Safety factor of 1.3 is used by coal industry in the United States [34] as a reference value for decision-making about the safety of the pillar system. Same safety factor is used here. Software gives safety factor values for different loading situations [34]: 1. Development loading - The loading on the pillar system before any longwall retreat mining. It is equal to the tributary area load. 2. Headgate loading - The pillar loading adjacent to headgate corner of the longwall face, which is equal to the development load plus the first front abutment. 46

48 3. Bleeder loading - The loading on a pillar system adjacent to a mined-out panel, which equals the development load plus the first side abutment. 4. Tailgate loading - The loading on a double-use gate entry system when it is adjacent to the tailgate corner of the longwall face, equal to the development load plus the first side abutment plus the second front abutment. 5. Isolated loading - The loading on a pillar system located between two mined-out panels, equal to the development load plus two side abutments. Two different gateroad development systems are compared in this paper, two-entry system and three-entry system (see chapter 2.4). Two-entry system needs less development time and resources, also less oil shale is left underground. Three-entry system is easier to ventilate, has more room for travel and increases gateroad stability [30]. Purpose is to find out which method would suit better for Narva oil shale mine. Results are presented in chapter

49 4. RESULTS Results from different analysis methods used in this thesis paper are presented in this chapter. Results for both parts of the work (highwall slope- and chain pillar stability analysis) are presented here. 4.1 Numerical modelling results for the highwall slope stability The results and explanations of the 10-stage FEM modelling by single stages are presented in Figure 20 and Figure 21. More thorough explanation can be found further down. In Stage 1 the length of longwall (i.e. length of the mined section) is 7 meters, after that in each stage the longwall advances by 5 meters in Stage 2 it is 12 m, in Stage 3 17 m, etc. Stage 10 corresponds to longwall length of 52 m. The color-code in the figures corresponds to the percentage of yielded elements. The dark blue color shows the areas where 0 % of the elements have yielded while red color indicates zones with 100 % of yielded elements. These zones indicate whether the roof is caving or not unstable or stable. In Figure 19 the Sigma 1 stress plots are shown for both numerical models FLAC and RS 2. It is shown here, because of comparing different program models with each other to verifty the proposed geotechnical model. Both of the models look similar when taking into account the major in-plane principal stress contour plots. It can be seen that the highest stresses are at the bottom of the model, which is expected. Also, the magnitude of stresses is similar in these models (around 1.8 MPa is the highest stresses). With the FLAC model the input stresses were calculated and with RS 2 the model calculated the stresses itself as the gravity stress option was used with the actual ground surface option. It can be concluded that these two models coincide nicely with each other. 48

50 Figure 19. FLAC and RS 2 initial model' Sigma1 stress plots. The modelling of Narva experimental panel shows that the first immediate roof weighting distance is only about 7 meters (Figure 20, Stage 1). According to the modelling results the non-commercial oil-shale (immediate roof, Layer 8) caves periodically almost as supports are moved. The limestone roof (total thickness of 7 m) caves periodically, depending on the thickness of the caving layer(s), after 6-9 m of temporary weighting (Figure 20, Stage 3). The thickness of caved layers may vary depending if the 7m thick roof is split into two blocks or not. 49

51 The modelling showed that the roof is caving in at least two stages firstly, the immediate roof (limestone and kukersite layers) caves and after that with advancing longwall length the main roof (limestone) caves in two stages. Advancing from Stage 1 to 2 the length of mined longwall advances from 7 m to 12 m and the percentage of yielded elements area increases towards the free surface, indicating a high probability of immediate roof caving. When the mined length reaches 27 m (Stage 5) the immediate and the main roof have already caved as suggested by the fact that most of the elements in that area have yielded. The barrier pillar up to this point is undisturbed no yielded elements. The yielding of the barrier pillar starts in Stage 6 with the mining face being 70 m from the highwall. In Stage 7 (65 m from the highwall) 2/3 of the barrier pillar elements are yielding. In Stage 8 (60 m from the highwall) the barrier pillar has yielded in its entire length. From these results it can be determined that the minimum barrier pillar length should be at least meters. To get the minimum length for the barrier pillar lets look at some more modelling results. Figure 20. Yielded elements in Stages 1-5 with the first figure showing the initial condition. The mined longwall length is from 7 to 27 meters. Color-code corresponds to the percentage of the yielded elements. Color-code: blue: 0%, red: 100%. 50

52 Figure 21. Yielded elements in Stages The mined longwall length is from 27 to 52 meters. Color-code: blue: 0%, red: 100%. The total rock overburden above the mine is about 16 meters. Figure 20 Stages 2 and 3 show a 4-9 meter thick rock massive deformation is following the roof caving. As a result, the rock massive above the oil shale seam is broken even before the longwall face advances (Figure 20 and Figure 21). It also indicates that the thin rock overburden is literally broken due to tension formed in the massive as mined panel advances and roof caves. The broken zone in the overburden rocks reaches about 20 meters ahead the working face towards the unmined area (Figure 20 Stage 5). 51

53 Figure 22. Total displacement curves of the barrier pillar from the highwall slope to 50 m inside the commercial oil shale layer. Figure 22 demonstrates the model output regarding the barrier pillar displacement by stages by total displacement vs distance in the barrier pillar curves. The distance is 50 m from the highwall slope inside the rock mass. In Stages 1-8 the displacements are quite marginal, close to zero, but beyond Stage 8 (60 m from the highwall) the deformations are non-linear indicating towards the failure of the barrier pillar. 52

54 Figure 23. Yielded elements curves of barrier pillar (50 m distance) Stages Figure 23 displays the results of yielded elements in the barrier pillar along the distance of 50 m from the highwall slope inside the commercial oil shale layer. From these results it can be determined that in Stage 7 (65 m from the highwall) the barrier pillar is starting to yield around 18 m inside the rock mass, but only 35 % of the elements in that area are yielding, so it in all probability still can take some load. From 21 m distance the yielded elements increase to 65 % and quite rapidly to 100 %. So, all in all, Stage 7 (65 m barrier pillar) is the minimum pillar length that can be used to provide adequate protection to the highwall. Stages 8-10 (60-50 m barrier pillar) are yielding from the highwall slope, so unstable. Figure 24 shows the total displacement contour plots in various stages. At first, the initial state is stable as the total displacement contour plot shows a 1 mm total displacement. In Stage 1 the displacement in the roof increase to 21 cm (Figure 24). It escalates up to 51 cm in Stage 5 without any yielding in the barrier pillar, but in Stage 6 (68 cm of displacement) some areas in the barrier pillar start to yield (Figure 21). The displacement will rise from Stage 5 51 cm to 160 cm in Stage 8 (Figure 25), which indicates a massive failure of the barrier pillar. After analyzing the total displacement contour plots it can be determined that 65 m (Stage 7) long barrier pillar is the minimum length for it, most probably 70 m will result with a more stable barrier pillar. 53

55 Figure 24. Initial state and Stage 1 total displacement contour plots (red dots indicating yielding elements). 54

56 Figure 25. Stage 5 and 8 total displacement contour plots (red dots indicating yielding elements). Figure 26 and Figure 27 show the plasticity indicators results from the FLAC software. From those one can conclude that the immediate roof starts to cave in Stage 1, when the mined longwall length is 7 meters. The highwall slope damaged zone is indicating failure in all stages, this can be because its assumed properties can be underestimated. The damaged zone might have stronger rock than estimated in the modelling, because it is stable in the site. When the mined longwall advances towards the higwall a large area around the longwall is indicating a shearing failure. From those different stage models one can definitely see that 55

57 the overburden is caving above the mined out layer. The FLAC results are displaying similarity to RS 2 plasticity plots. Beginning from Stage 5 the barrier pillar is starting to show some yielding. From Stage 8 the pillar gets unstable, because of the percentage of yielded elements increase towards the highwall. Half of the length of the barrier pillar has yielded in that stage. In Stage 9 and 10 the barrier pillar is yielding in its entire length. 65 (Stage 7) or 70 (Stage 6) meter barrier pillar will quarantee a stable highwall based on the FLAC plasticity indicator plots. Figure 26. Plastiticity indicators in Stages 1-5 with the first figure showing the initial condition. The mined longwall length is from 7 to 27 meters. Plasticity indicator: orange shear (neg); green tension (neg); purple shear (pos); yellow tension (pos). 56

58 Figure 27. Plasticity indicators in Stages The mined longwall length is from 27 to 52 meters. Plasticity indicator: orange shear (neg); green tension (neg); purple shear (pos); yellow tension (pos). Figure 28, Figure 29 and Figure 30 display the x-displacement contour plots. They are from Stage 6 to 8, so the barrier pillar length will be from 70 to 60 meters. The plots show horizontal movement of the elements of the model. One can see that in Stage 6 (Figure 28) the red area indicates zero displacement, while other areas around it are moving slightly. Furthermore, in Stage 7 (Figure 29) the orange area above and the barrier pillar itself is indicating zero displacement. In Stage 8, when the remaining barrier pillar length is 60 meters, Figure 30 indicates around 1-meter displacement of the area surrounding the pillar. From the x-displacement contour plots, it can be determined that the 65-meter-long barrier pillar will quarantee an adequate stability for the highwall slope. 57

59 Figure 28. X-displacement contours in Stage 6. Figure 29. X-displacement contours in Stage 7. 58

60 Figure 30. X-displacement contours in Stage Results from ALPS (Chain pillar stability) Graphs are based on excel tables presented in appendices 5 and 6. Figure 31 shows safety factors for a 3-entry system with a pillar length of 4 meters. Only development loading meets the required safety factor of 1.3 when width of the pillar is over 7 meters, all other loading conditions are under it. Isolated loading stays under SF of 1.0 even if width of the pillar is 8 meters. Pillars are unstable. 59

61 Safety factor Safety factor Chain pillar SF, Pillar length = 4 m, 3-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 31. Chain pillar safety factor, pillar length = 4 m, three-entry system. Figure 32 shows safety factors for a 2-entry system with a pillar length of 4 meters. Only development loading meets the required safety factor of 1.3, when width of the pillar is over 7 meters, all other loading conditions are under it. Isolated loading stays under SF of 1.0 even if width of the pillar is 8 meters. Pillars are unstable. There is no considerable difference between 2-entry and 3-entry systems in this case. Chain pillar SF, Pillar length = 4 m, 2-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 32. Chain pillar safety factor, pillar length = 4 m, two-entry system. Figure 33 shows safety factors for a 3-entry system with a pillar length of 5 meters. Only development-, headgate and bleeder loadings meet the required safety factor of 1.3 when width of the pillar is over 7 meters, all other loading conditions are under it. Isolated loading meets SF of 1.0 when pillar width is slightly over 6 meter, but requires width over 8 m in order to meet SF of

62 Safety factor Safety factor Chain pillar SF, Pillar length = 5 m, 3-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 33. Chain pillar safety factor, pillar length = 5 m, three-entry system. Figure 34 shows safety factors for a 2-entry system with a pillar length of 5 meters. Only development loading and headgate loading meet the required safety factor of 1.3 within the set interval of chain pillar size, all other loading conditions are under it. Isolated loading stays under SF of 1.0 even if width of the pillar is 8 meters. Pillars are unstable. Compared to 3 entry system it gives inferior results. Chain pillar SF, Pillar length = 5 m, 2-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 34. Chain pillar safety factor, pillar length = 5 m, two-entry system. Figure 35 shows safety factors for a 3-entry system with a pillar length of 6 meters. All loading conditions meet the safety factor of 1.3 when width of the pillar is around 6 meters. Pillar with a size of 6x6 meters would be safe enough for stability concerns and could be used for longwall mining in Narva deposit. 61

63 Safety factor Safety factor Chain pillar SF, Pillar length = 6 m, 3-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 35. Chain pillar safety factor, pillar length = 6 m, three-entry system. Figure 36 shows safety factors for a 2-entry system with a pillar length of 6 meters. Development-, headgate and bleeder loadings meet the required safety factor of 1.3 when width of the pillar is 6 meters. Tailgate loading is slightly above safety factor of 1.0 and isolated loading under it when pillar width is 6 meters. Only isolated loading can be deemed as unstable. 3-entry system on the other hand gave much better SF results under same pillar dimensions. Chain pillar SF, Pillar length = 6 m, 2-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 36. Chain pillar safety factor, pillar length = 6 m, two-entry system. Figure 37 shows safety factors for a 3-entry system with a pillar length of 7 meters. All loading conditions meet the required safety factor of 1.3 under pillar width around 6 meters. 3- entry system with pillar length of 6 meters is preferable, it stays stable under same conditions with same pillar width. 62

64 Safety factor Safety factor Chain pillar SF, Pillar length = 7 m, 3-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 37. Chain pillar safety factor, pillar length = 7 m, three-entry system. Figure 38 shows safety factors for a 2-entry system with a pillar length of 7 meters. All loading conditions meet the required safety factor of 1.3 when pillar width is slightly over 7 meters. With pillar width of 6 meters tailgate- and isolated loading are both equal or slightly above safety factor of 1.0. This system should be relatively stable. Chain pillar SF, Pillar length = 7 m, 2-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 38. Chain pillar safety factor, pillar length = 7 m, two-entry system. Figure 39 shows safety factors for a 3-entry system with a pillar length of 8 meters. All loading conditions meet the required safety factor of 1.3 under pillar width slightly under 6 meters. 3-entry system with pillar length of 6 meters is preferable, it stays stable under same conditions with similar pillar width. 63

65 Safety factor Safety factor Chain pillar SF, Pillar length = 8 m, 3-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 39. Chain pillar safety factor, pillar length = 8 m, three-entry system. Figure 40 shows safety factors for a 2-entry system with a pillar length of 8 meters. All loading conditions meet the required safety factor of 1.3 when pillar width is 7 meters. With pillar width of 6 meters tailgate- and isolated loading are both slightly above safety factor of 1.0. This system should be relatively stable but inferior to the 2-entry verion with pillar length of 7 meters as the safety factors are very similar. Chain pillar SF, Pillar length = 8 m, 2-entry Width [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 40. Chain pillar safety factor, pillar length = 8 m, two-entry system. Figure 41 shows chain pillar safety factor for three-entry system relative to length when width of the pillar is fixed at 3 meters. Compared to previous graphs it can be easily determined that when width is low, even very high lengths valued won t suffice in order to meet the required safety factor value of

66 Safety factor Safety factor Chain pillar SF, Pillar width = 3 m, 3-entry Length [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 41. Chain pillar safety factor, pillar width = 3 m, three-entry system. Figure 42 shows chain pillar safety factor for two-entry system relative to length when width of the pillar is fixed at 3 meters. Similar to the 3-entry system it also does not meet the safety factor requirement even if length of the pillar is increased greately. Chain pillar SF, Pillar width = 3 m, 2-entry Length [m] Developement loading Headgate Loading Bleeder loading Tailgate loading Isolated loading Figure 42. Chain pillar safety factor, pillar width = 3 m, two-entry system. Table 15 shows chain pillar dimensions, which meet the required safety factor requirement, in relationship to area of the pillar. As length of the pillar increases over the lowest suitable value, width reduces, but area of the pillar also increases. This means that if we increase dimensions over a certain optimal value then size of the pillar increases but the stability remains similar. 65

67 Table 15. Chain pillar dimensions in relationship to area of the pillar. 3-entry system (SF=1.3) Dimensions (LxW) Area [m] [m 2 ] 6x6 36 7x5.8 40,6 8x5.4 43,2 2-entry system (SF=1.3) 7x6.7 46,9 8x

68 5. DISCUSSIONS AND RECOMMENDATIONS Discussions and recommendations of this thesis paper are presented here. Chain pillar and highwall slope stability analysis are handled separately. 5.1 Higwall slope stability analysis A properly designed barrier pillar plays a significant role in the stability of the highwall slope, since it is an only safeguard that remains between the final longwall position and the highwall, which is subject to redistributed ground stresses. Material is strained to magnitudes unlike those typically measured in open-cut mining. Mine personnel and equipment may become exposed to the unstable highwall rockfall hazards, highlighting the importance of understanding the mechanism and implementing appropriate controls. [38] The intention of numerical modelling was to determine the behavior of the material surrounding the pillar area, rather than absolute values. In introducing residual values through the assumption of plastic behaviour, the finite element model for this scenario has heavily over exaggerated the magnitude of deformation in the caving zone compared to that known, and realistic in the actual environment [38]. To begin with, the results from the finite element program RS 2 indicate that a 65-meterlong barrier pillar will yield a stable highwall. Firstly, the yielded elements plots are demonstrating that the minimum length for the barrier pillar might be 65 meters. Secondly, the yielded elements curves are also exhibiting results that 65 meters long barrier pillar is the minimum length for it. Thirdly, the total displacement plots confirm the previous results: minimum length for the barrier pillar that will still provide adequate protection for the highwall is 65 meters. Furthermore, the results from the finite difference elements program FLAC are corroborating the results from RS 2. Firstly, after analyzing the plasticity indicator plots the minimum length for the barrier pillar that will quarantee a stable highwall is 65 meters. Secondly, the x-displacement (horizontal movement) plots are manifesting that the 65-meterlong barrier pillar will quarantee an adequate stability for the highwall. 67

69 Both numerical models are denoting towards the 65-meter-long barrier pillar. It can be safe to say after analyzing the results from both numerical programs that the minimum recommended length for the barrier pillar is 65 meters. However, the shrinking and fracturing of the barrier pillar walls can be expected after the walls are exposed. When the barrier pillar wall collapses (decrease in the barrier piller width/length), this leads to increasing load per area of the barrier pillar. The oil shale strength inside the barrier pillar after the longwall roof failure is also decreased due to tension and shear forces, which in turn is causing the formation of new discontinuities in the oil shale seam. These processes might lead to the self-collapsing of some parts of the barrier pillar in time. This phenomenon has to be taken into account before designing the barrier pillar dimensions. The following recommendations can be used in order to make the highwall slope more stable or if for example shorter length for the barrier pillar than suggested above will be used: artificially supporting the highwall slope may be a solution. For example, reinforcement such as bolts, cables, mesh, and shotcrete to support the rock mass. The use of such supports can be very expensive. A careful study of the geologic structures must be performed to select the proper reinforcement (i.e. length of bolts or cables, thickness of shotcrete, etc.). Bolts that are too short will do little to prevent slope stability problems from continuing; [39] if a slope continues to fail, and supporting the slope is not a feasible alternative, steps need to be taken to remove the hazard. Often, flattening the slope to a more favourable angle with respect to the local geology will solve the problem [39]. That might be the last solution, because of high costs or if the need to remove the barrier pillar is high; since water pressure creates slope stability problems, dewatering using horizontal or vertical wells is a powerful means of controlling slope behaviour and minimizing hazards. [39] 68

70 5.2 Chain pillar stability analysis In chapter 4.2 many scenarios regarding safety factors in Narva longwall mine were developed and presented. Different width, length and entry systems were analyzed in order to determine the most optimal design parameters for a future mine. Corresponding geological and rock mechanical data for Narva deposit were used. Based on experience from room and pillar mining in Estonia it was estimated that optimal parameters for a chain pillar in Narva should be between 4-8 meters. This estimation was tested in this thesis paper and found to be true. If safety factor of 1.3 is taken into account, then no analyzed cases have sufficient stability when pillar length or width is between 4-5 meters. Based on previous information, one can determine that anything lower than that would surely give even worse results. Optimal parameters were found to be around 6 meters which means that pillar parameters over that would be waste of resources because gateways would stay safe in both cases. During the making of this thesis paper question aroused that what if width of the pillar is reduced further, below the set parameters, and length is enhanced greatly. It is determined that this would not give good results as high length, with 3-meter width, would not meet the set 1.3 safety factor value. This applies for both two-entry and three-entry systems (see Figure 41 and Figure 42). It can be determined that increasing length increases the safety factor, but below optimal value it starts to play relatively marginal role. Optimal pillar length differs between 3-entry system and 2-entry system. For a 3-entry system, pillar length of 6 meters and width of 6 meters meets the safety factor requirement of 1.3 for all loading cases. Further increase in either length or width of the pillar gives marginal improvements while area of the pillar increases more rapidly (see Table 15). For a 2-entry system pillar parameters of 7x7 meters are determined to be optimal as lower values don t meet the safety factor requirement and higher values give marginal improvement (see Table 15). Based only on the stability standpoint, then 3-entry system is best and 6-meter pillar width is suggested (see Figure 35). Compared to 2-entry system this means more development work and less (although marginal) resources extracted as oil shale in pillars are left underground. Best alternative from 2-entry system would be 7x7 meter pillar which also meets the safety factor requirement of 1.3. This on the other hand would be harder to ventilate based 69

71 on the coal mining experience [30], maingate needs to be the main intake for fresh air and tailgate is the outlet for contaminated air. These are empirical results that need further field investigation to determine the sufficient safety factor values. It could be possible, that safety factor less than 1.3 is suitable for safe oil shale mining with a longwall method, meaning that lower dimensions for chain pillars can be used. Also, much longer pillars may be used to reduce the development time and cost, resulting in slightly lowered width value for pillars. 70

72 REFERENCES 1. Rell, M. Kupts, M. (2004). SA Poliitikauuringute Keskus Praxis. Põlevkivitööstuse mõju demograafilistele arengutele kuni aastani Tallinn. 2. Härm, M. Põlevkivi rahvuslik rikkus või õnnetus? [WWW] ( ). 3. TPÜ Ökoloogia instituut. (2003). OÜ VKG Oil poolt kavandatava Ojamaa põlevkivikaevanduse rajamise ja põlevkivi kaevandamisega kaasneva keskkonnmõju hindamine. Jõhvi. 4. Adamson, A. Mäeinsituut. (2005). Eesti põlevkiviressursi kasutamissuundade riikliku strateegia aastani 2020 alusuuringud. Tallinn: TTÜ Mäeinstituut. 5. Estonian Environmental Information Centre (EEIC). (1994). Estonian Mineral Resources. Tallinn. 6. Aaloe, A. (2006) Kukersiit Eesti põlevkivi. Tallinn: MTÜ GEOGuide Baltoscandia [WWW] ( ). 7. Cane, R.F. (1976). The origin and formation of oil shale. Oil Shale. Amsterdam: Elsevier. 8. IEA (2010). World Energy Outlook Paris: OECD. 9. Reinsalu, E. (2011) Eesti mäendus. Tallinn, Eesti: Tallinna Tehnikaülikooli Kirjastus. 10. Dyni, J. R. (2006) Geology and Resources of some World Oil-Shale Deposits. Virginia: U.S. Geological Survey. 11. Kattai, V. Saadre, T. Savitski, L. (2000) Eesti Põlevkivi. Tallinn: Eesti Geoloogiakeskus. 12. AS Maves. OÜ Inseneribüroo STEIGER. (2014). Enefit Kaevandused AS kaevandamislubade KMIN-073, KMIN-046, KMIN-074 ja KMIN-087 muutmisega kaasneva keskkonnamõju hindamise (KMH) programm. Tallinn. 13. Veljo, A. (2014). Taotlus maavara kaevandamise lubade muutmiseks ja kaevandamise meetodi lisamiseks [WWW] ( ). 14. Estonian Land Board: Geoportal. Mining shares detailed information. 71

73 15. AS Eesti Põlevkivi Tootmisosakond. (2005). Narva karjääri mäetöödega rikutud maakorrastamise projekt. Jõhvi. 16. Tohver, T. (2011). Põlevkivi kaevandamis- ja rikastamisjääkide kasutamine. Tallinn: TUT Press. 17. Viil, A. (2011). Põlevkivikarjääride (kaevandatud alade) rekultiveerimine. Valgma, I. (Toim.). Kaevandamine ja vesi ( ). Tallinn: TTÜ Mäeinstituut. 18. Ratas, M. (2001). Narva Karjääerivälja puuraugud. Tallinn: Eesti Geoloogiakeskus. 19. Simons, N., Menzies, B. & Matthews, M. (2001). A Short Course in Soil and Rock Slope Engineering. ICE Publishing. 20. Hoek, E. (2006). Practical Rock Engineering. Online Handbook. Available at Tavast, E., Raukas, A. (1982). Eesti aluspõhja reljeef. Tallinn, Valgus. 22. Sokman, K., Kattai, V., Vaher, R. & Sõstra, Ü.J. (2008). Influence of tectonic dislocations on oil shale mining in the Estonia deposit. Oil Shale, 25 (2SI), pp Field work observations. 24. Nikitin, O. (2013). Summary of underground mining experiments with roof caving from different mines and Rock Input parameters. The summary was provided by the mining company. 25. Gercek, H. (2007). Poisson s ratio values for rocks. International Journal of Rock Mechanics and Mining Science, 44 (1), Sheory, P.R. (1994). A theory for in situ stresses in isotropic and transversely isotropic rock. Int. J. Rock Mech. Min. Sci. & Geomech. Abstr. 31 (1), Zoback, M. L. (1992). First- and second-order patterns of stress in the lithosphere: The World Stress Map Project. J. Geophys. Res. 97 (B8), Palchic, V. (2003). Formation of fractured zones in overburden due to longwall mining. Environmental Geology, 44, Raukas, A., Tavast, E., Vaher, R. (2007). Vasavere ancient valley, its morphology, genesis and importance in the economy of North-East Estonia. Baltica, Vol. 20 (1-2), Vilnius. 30. Peng, S.S. (2006). Longwall mining, 2 nd edition. USA: West Virginia University. 72

74 31. Tätsep A., Toomik A., Liblik Valdo. (2003). OÜ VKG Aidu Oil poolt kavandatava Ojamaa põlevkivikaevanduse rajamise ja põlevkivi kaevandamisega kaasneva keskkonnamõju hindamine. Jõhvi: TPÜ Ökoloogia Instituut. 32. Oraee, K., Hosseini, N., Gholinejad, M. (2010). Optimization of Chain Pillar Design in Longwall Mining Method. The 29 th International Conference on Ground Control in Mining. 33. Mark, C., Bieniawski, Z.T. (1986). An Empirical Method For the Design Of Chain Pillars For Longwall Mining. American Rock Mechanics Association. 34. The National Institute of Occupational Safety and Health. Mining Product ALPS Analysis of Longwall Pillar Stability. [WWW] ( ). 35. Galvin, J.M. (2016). Ground Engineering Principals and Practices for Underground Coal Mining. ACARP. 36. TTÜ Geoloogia Instituut. (2007). Tootuskihind ja lasund Põhja Kiviõli põlevkivikarjääris. Tallinn [WWW] ( ). 37. Aruküla, H., Eigo, L., Joosep, E., Reinsalu, E. (1980). Puur- ja lõhketööd: Käsiraamat. Tallinn. 38. Clarkson, L. (2016). Effect of punch longwall retreat on highwall stability. Proceedings of the 16th Coal Operators' Conference, Mining Engineering, University of Wollongong, February 2016, Girard, J. M. (2001). Assessing and monitoring open pit mine highwalls. Proceedings of the 32nd Annual Institute on Mining Health, Safety and Research, Salt Lake City, Utah, August 5-7, Jenkins FM, Langton J, McCarter MK, Rowe B, eds. Salt Lake City, UT: University of Utah,

75 APPENDICES Appendix 1. Geological crossection of Northern-Kiviõli oil shale quarry [36] 74