Quality Assessment of Backfill Performance for an Underground Iron Mine in Turkey

Ground Support 2016 A.A. Editor and B. Editor (eds) ISBN 978-0-98709xx-x-x Quality Assessment of Backfill Performance for an Underground Iron Mine in Turkey A.G. YARDIMCI, Middle East Technical University, Turkey L. TUTLUOGLU, Middle East Technical University, Turkey C. KARPUZ, Middle East Technical University, Turkey H. OZTURK, Middle East Technical University, Turkey D. GUNER, Middle East Technical University, Turkey Abstract A hybrid underground mining method is selected for Karacat iron orebody located mid-south east of Turkey. For production levels at higher elevations above 1497, sublevel stoping with stope dimensions of 10 x 25 m is proposed. Below this level, production is planned to proceed with 5x5m cut & fill stopes. Production sequence is arranged for primary and secondary stopes in the sublevels of upper levels. Rock mass properties for the orebody and host rock units were characterized by geotechnical investigations. Results of laboratory experiments and classification work provided input data for numerical models. Considering technical and operational limitations and orebody geometry, various backfill scenarios were modelled. Alternative fill characteristics were investigated to sustain global stability. A relatively high quality fill with an unconfined compressive strength (UCS) of 1.5 MPa for primary stopes, and a lower quality backfill with a UCS of 0.75 MPa for secondary stopes were computed to be sufficient for the stability. For verification of the backfill quality, core samples collected from a field practice were tested in the laboratory. Results of uniaxial compression tests illustrated wide scatter. Observations pointed out that non-uniform particle size distribution with extremely large particles caused insufficient cementation in the backfill. An aggregate mixture with -30 mm particle size distribution and 6-8% cement ratio was recommended for the primary fill. To improve the performance of the filling in practice, waste rock to be used in filling practice was suggested to be processed by a crusher to assure the proper particle size distribution before mixing with cement. 1 Introduction Large volumes of excavated spaces are left behind as a natural result of underground mining activities. Disturbance of initial stress field in the rock mass due to excavation may lead to global instability problems. Deformation in rock mass provides relaxation for the rock mass stress distribution compared to the induced stresses right after the excavation. Deformations may exceed the limits of elastic rock behavior as a consequence of unsupported volumes of relatively large excavations. Plastic failures fill these voids at the expense of underground mine global stability or surface subsidence in the topography. Such occurrences may slow down or completely halt ore production. Although caving methods are attractive in terms of low production cost, safety concerns may dominate the underground mining method selection. Therefore, supported mining methods might be the first choices of mining companies when global mine safety becomes questionable. There are two major performance expectations from the backfilling practice in this underground mine. Backfilling is expected to guarantee the stability of individual local stopes during production at different levels. A reliable performance is to be provided by the fill support system during the production Ground Support 2016, Luleå, Sweden 1

Ground Support 2016 proceedings A.F. Surname and J.S.A. Surname throughout the life of the mine. Second expected performance from the filling practice is to maintain a globally safe structural state for the whole orebody, as excavation operations come close to the final stages of stoping and production work in the mine. Commonly used combinations for backfills of different type include dry sand and rock fills, non-cemented and cemented hydraulic fills, cemented waste-rock fills, paste fills of processing plants, pneumatic fills and fluent fills (Hambley, 2011). Considering the composition, application procedure, and particle size, each of them has advantages and disadvantages. The simplest and mostly preferred choice is to employ dry sand and rock fill. For this, waste rock of the mine from the completed stopes is directly applied with no need to reduce particle size and no cement mixture. Transportation is the single cost item. Low stiffness of this type of backfill is the main disadvantage. Another widely used filling practice includes cemented rock fills. Low strength problem of rock fills is treated by adding a cement mixture. Particle size distribution dramatically affects the quality and performance of cemented rock fills. Backfill composition with coarse size particles may lack planned strength due to cementation problems as a result of low particle surface area in contact with the cement mixture. For Karacat region, 3D orebody model and DTM (digital terrain model) of topographical surface are identified and extracted by using the field data in the related simulation program. Massive type iron orebody is surrounded by smaller orebodies. For planned underground iron mine in Karacat-Kayseri/Turkey region, backfill design following the stoping sequence is assessed. Results of rock mass quality characterization work and laboratory experiments are used to estimate the input mass parameters to be used for the stability analyses by numerical modelling. Economical, practical and structural considerations revealed two backfill types as appropriate for primary and secondary stopes. A hybrid mining method with combination of sublevel stoping and cut & fill stoping is decided to be applicable stably with the recommended backfill characteristics. Unconfined compressive strength tests were conducted on core samples of filling material (containing particles above 30mm and 6-8% cement mixture) collected from field practice of early filling operations. Purpose was to check whether the planned quality of the backfilling support system was being satisfied as recommended in planning. 2 Mine Location and General Descriptions Karacat underground mine is located right below the open pit (see Figure 1). Apparently, open pit slope stability will be risky if proper filling design is not applied in underground mining operation. Three critical items of underground mine design are investigated: stope dimensioning, pillar design and backfill design. As part of the overall project, production stope dimensioning is first conducted. Later, pillar stability work is carried out for the multiple stope production operations at different levels of mine. Finally, economically optimum backfill alternatives are assessed from the point of local and global structural stability by numerical modeling. 2.1 Location Karacat mine is located in Yahyali district of Kayseri / Turkey. A satellite view of the mine, plan view from the 3D model showing the orebody, orebody dimensions and production levels from the cross section view can be seen in Figure 1. Besides from the Karacat mine, there are four active mines: two open pit and two underground mines in the area. 2 Ground Support 2016, Luleå, Sweden

Proceedings Section/Chapter Front view A B B A Plan view Front view Figure 1 Location of the mine 2.2 Geology Geology of Karacat iron orebody was studied by Tiringa (2009) in the scope of a Master of Science work. The Geyikdag unit was described to be located in the Taurid Tectonic Belt hosting the Karacat iron orebody. The orebody was reported to be surrounded by Emirgazi (Precambrian), Zabuk (Lower Cambrian), Değirmentaş (Middle Cambrian) and Armutludere (Ordovisian) formations. Hematite and goethite are the major ore minerals which are believed to have originated as a product of siderite alteration. The ore body and country rocks interrelation (Zabuk formation, Değirmentaş formation and Armutludere formation) can be stated to be controlled by tectonism. Surface reaction mechanism and karstification processes are the result of Post-mineralization faults. Altered siderite and iron minerals transform into limonite and goethite predominated by atmospheric conditions where surface reaction mechanisms are active. Products of the mineralization process mentioned above are exploited and served to industry as raw material. Karacat Iron ore deposit can be described as a deformed deposit occurred by flow of hydrothermal fluids from Precambrian aged primer iron deposits. 3 Geotechnical Studies 2D plane strain numerical modelling is carried out for pillar stability and backfill design. Input parameters of numerical models are decided based on the field studies and laboratory tests. Rock mechanics tests and empirical classifications are conducted by mainly logging and testing on drill core samples. Ground Support 2016, Luleå, Sweden 3

Ground Support 2016 proceedings A.F. Surname and J.S.A. Surname 3.1 Field Studies In the scope of a field trip, rock mass characterization, discontinuity mapping and sample selection for laboratory testing tasks were completed. It was observed that the main ramp excavation face was completed up to the 1497 level. Also, some preliminary stopes were driven. Calcschist and limestone were observed dominantly both in the hanging wall and foot wall. Ore minerals were seen to be in the form of hematite, goethite, siderite and alterations like limonite, which appeared relatively weaker at the first sight. Empirical classification oriented characterization of rock mass was done both on samples in core boxes and at excavation faces of outcropping parts. Appropriate specimens for laboratory testing were collected. 3.2 Laboratory Tests In METU (Middle East Technical University) Rock Mechanics Laboratory, uniaxial compressive strength tests, triaxial compression tests, static deformability tests, dry unit weight tests and indirect tensile strength (Brazilian) tests according to ISRM suggested methods are carried out on intact rock samples obtained from borehole core specimens. Summary of laboratory test results are presented in Table 1. Table 1 Laboratory test results of intact rock specimens Test Ore Hanging Wall - Foot Wall Unconfined compressive strength, UCS (MPa) 45.5 89.8 Young modulus, E (GPa) 12.6 21.1 Poisson s ratio, ν (MPa) 0.17 0.11 Cohesion, c (MPa) 11.4 10.1 Internal friction angle, ( ) 51.3 53.5 Tensile strength, σ t (MPa) 4.8 9.8 Unit weight, γ (kn/m 3 ) 29.4 27.1 3.3 Rock Mass Characterization The GSI system was used to characterize rock mass exposures on the excavation faces in the field. RMR 89 (Bieniawski, 1989) and Q system (Barton, 1974) are used to define rock mass quality through a depth of 30 m in the hanging wall and a depth of 20 m depth in the footwall. The reason for concentrating on these critical depth bands in the hanging and footwall can be explained by their effectiveness in the local and global stability of the mine. Shear zones located within the orebody are about 3-5 m thick. Shear zones in the hanging wall and footwall are 1-3 m thick. All of these zones show repetitions at every 20 to 30 m. Average ratings and quality descriptions can be seen in Table 2. Table 2 Average rock mass quality scores with respect to GSI, RMR and Q systems Rock Unit GSI RMR 89 RMR Quality Desc. Q Q Quality Desc. Hanging Wall 53 58 Fair 4.2 Fair Ore 45 49 Fair 3.4 Poor Footwall 40 45 Fair 1.0 Poor 4 Ground Support 2016, Luleå, Sweden

Proceedings Section/Chapter 4 Backfill Design and Quality Assessment Underground mine stability is critical even after the exploitation is completed. Surface subsidence triggered by underground activities can lead to serious risks for the overlying abandoned pit slopes. This instability in turn can present a major problem for underground workings. In addition to protecting the neighboring local stopes, backfilling has become a standard cycle of underground mining in spite of its increasing effect on the mining cost. Numerical models confirmed that production below 1497 level of the mine with stoping techniques dramatically affected global mine stability. Therefore, a hybrid underground mining method was considered and analyzed. Sublevel stoping method was planned to be applied above 1497 level. Below this level, ore would be extracted by cut & fill method. For primary and secondary stopes, two backfill alternatives are studied. Effect of backfill quality on stability of local orebody pillars is analyzed as the first issue. Towards the end of the production in a mine level, a few stopes which are not extracted yet, behave as only supporting orebody pillars. They are overloaded and contain the risk of pillar burst due to high stress concentrations. FEM models are constructed to analyze this risk. At the final stage, completely filled underground mine stability is analyzed for checking the global stability state. 4.1 Modeling Entries for Stability Analyses Numerical analyses are carried out in Phase2 Finite Element Method (FEM) software with the assumption of plane strain condition. FEM models are prepared on two critical cross sections across the orebody using 3D topographical surface and orebody solid model (Figure 1). Rock mass properties are estimated by using GSI values and intact rock UCS values in ROCKDATA software. Output is a fitted curve for the Hoek&Brown failure criterion given in equation (1). From the Generalized Hoek&Brown failure criterion (Hoek et. al., 2002,) it is possible to estimate equivalent Mohr-Coulomb parameters as cohesion and internal friction of rock mass. σ a 3 σ 1 = σ 3 + σ ci (m b + s) σ ci (1) Where m b, s, and a are well-known constants representing the different structural states of the rock mass. For plastic analyses, GSI scores representing residual states are calculated from equation (2). Using laboratory tests and regular GSI scores, residual strength parameters are calculated based on the residual GSI given in equation (2) by Cai et.al., 2007. GSI res = GSIe 0.0134GSI (2) Peak material properties represent the transition from elastic to plastic material behavior. It aims to examine initial failure in structural units. However, residual material properties focus on the state after failure with increasing deformations. For instance, plastic material volume increases after failure due to broken rock pieces and this leads to greater deformations at the excavation boundaries. Volume increase is controlled by dilation angle parameter. Rock mass dilation angle in modeling work is imposed based on the results of equation (3) suggested by Alejano et. al. (2009): ψ = (5GSI 125)φ/1000 (3) For the residual rock mass tensile strength, ratio of tensile to compressive strength obtained from the laboratory tests is adopted. This ratio is around σt /σci = 0.1 for the laboratory strength; so it is kept the same to estimate the ratio of rock mass strength entries in the models. Predicted rock mass tensile strengths based on Hoek&Brown criterion are too low and cause instabilities in model integrity. Ground Support 2016, Luleå, Sweden 5

Ground Support 2016 proceedings A.F. Surname and J.S.A. Surname In Table 3, rock mass parameters for peak and residual states of rock units, and contact zone are given. Although backfill alternatives with varying mechanical properties are investigated in the models, only the ones considered in final decision is presented in Table 3. Primary backfill strength (particle size below 30 mm and 6-8 % cement mixture) is suggested be around at least 1.5 MPa. Secondary backfill is estimated to be satisfactory with a strength which is about half of the primary fill. Table 3 Input rock mass properties for numerical analyses Rock mass units Ore Hanging Wall Foot Wall Peak Res. Peak Res. Peak Res. Contact Zone Primary Backfill Secondary Backfill GSI 45 25 53 26 40 23 25 - - Unit weight (kn/m 3 ) 29.4 27.1 27.1 27.1 20.0 20.0 Modulus of Elasticity (GPa) 2.81 7.71 3.36 0.75 0.60 0.30 Poisson s ratio 0.17 0.11 0.11 0.17 0.30 0.30 Mohr - Coulomb Parameters σ cmass (MPa) Tensile Strength (MPa) Internal friction angle ( ) Cohesion (MPa) Dilation angle ( ) 11.2 7.0 30.7 17.1 23.8 15.9 7.0 1.5 0.75 1.1 0.7 3.1 1.7 2.4 1.6 0.7 0.2 0.1 37 31 42 34 39 33 31 35 35 2.8 2.0 6.8 4.5 5.7 4.3 2.0 0.4 0.2 4 6 3 - - - 4.2 Results of Modeling Work Above the 1497 level, 10x25 m stopes are planned to be applied. Primary and secondary stopes are proposed to be filled with the appropriate backfill materials. Below the 1497 level, 5x5 m cut & fill stopes are planned to be located; and all the stopes here are suggested to be filled with primary backfill material. Effect of backfill material on local stability of stopes and on the global stability of the mine are analyzed in detail. Elastic models are generated to compute the safety factor of the pillars that are locally protecting the production stopes. Stress distribution in pre-failure state will be greater compared to the post failure of a pillar. Thus, elastic model that represents the most critical state (pre-failure state) works better to predict the pillar safety factor. Data points of related structural entries are presented in the model outputs along the model query lines across the pillars. Using the principal stresses of the model outputs, rock mass strength state is estimated by comparing cohesion and internal friction angle based failure criterion to the stress state given by model outputs. Equation (4) is for the computation of rock mass pillar strength at varying confinement stress throughout a typical pillar: 6 Ground Support 2016, Luleå, Sweden

Proceedings Section/Chapter σ 1,strength = σ cmass + qσ 3 (4) Where σ cmass is global strength and σ 1,strength is rock mass compressive strength under confinement. Here q is confinement stress factor that can be calculated from internal friction of rock mass. Pillar walls remain unconfined unless support is applied. Instead of using rock mass uniaxial compressive strength (σc) suggested by ROCKDATA processing, global strength (σcmass) is preferred for computing FOS. Global strength is preferred to be used in comparison to the major principal stress for pillar wall stability investigations around the underground excavations. Direct use of σ c is not advised, since it suggests values which are too conservative for pillar design applications. Instead, use of global strength in Equation (5) is recommended by Hoek&Brown (1997) for stability checks at pillar walls. (m b + 4s a(m b 8s)(m b 4 + s) a 1 ) σ cmass = σ ci 2(1 + a)(2 + a) (5) Inside pillars where confinement exists, σ 1, strength of Equation (4) is used to estimate factor of safety as in equation (6): FOS = σ 1,strength σ1 (6) In a stope & pillar type mining, acceptable factor of safety is around 1.3 for local and global stability of the mine. In this study, models are constructed with 2D plane strain assumption. With this assumption, pillars are expected to be loaded more compared to the 3D structural state. So, plane strain modeling remains on the safe side. Thus, any FOS greater than 1.2 for pillars of 2D modeling is predicted to indicate that those pillars stay sufficiently on the safe. Plastic analyses in modeling work yield failure types (like shear or tensile failures), failure zone and geometry. This way, it is possible to comment on the effectiveness of any support type depending on the failure mode. Between the orebody and hanging wall, a shear zone of around 3-5 m thick is detected from the drill core sample observations. This part is named as contact zone and orebody geometry is encapsulated by such a zone in the models. In order to check the validity of the calculated σ cmass, existing trial stopes are modelled and interpreted using back analysis method. Current failure zones in opening walls and model results are compared. In Figure 2 total displacement contours and failure zones in a plastic analysis can be seen. From the field investigations it is roughly estimated that depth of failure zones is 600 mm in walls and 10 mm in roof. Compared to the numerical model it is concluded that calculated rock mass properties are representative for the field. Ground Support 2016, Luleå, Sweden 7

Ground Support 2016 proceedings A.F. Surname and J.S.A. Surname Total Displacement (mm) 0 mm 3 mm 1517 Level Figure 2 Total displacements and failure zones around existing preliminary stopes In the numerical models, production sequence starts from the orebody below 1497. Cut & fill stopes with dimensions of 5x5m are planned. Maximum four stopes are produced at the same time and the safety distance between operating stopes is 30 m. Production starts from the lower level that is 1477. Cut & fill stopes are completely exploited and filled till 1497. Later, sublevel stoping method is applied above 1497 level. First, primary stopes are produced and filled with primary backfill. Filled primary stopes behave just as pillar while secondary stopes are produced. Secondary stopes that are completed are filled with secondary backfill material. Section view of the model showing the end of production sequence can be seen in Figure 3. A contact zone presenting the deformed layer on the orebody wall rock contact can also be seen in this model. Hanging wall Contact zone Primary Backfill Foot wall Figure 3 Model view showing the end of production layout Backfill stability is examined separately at mine levels below and above the level 1497. Firstly, primary backfill performance in cut & fill stopes below level 1497 during production is analyzed. End of production sequence below level 1497 is modelled next. Stability of primary stopes are investigated for conditions at during production and at end of production stages above level 1497. Last stage of stability assessment for mining for above the level 1497 is for the secondary stopes. Results of numerical models are processed for during production and end of production stages here. 8 Ground Support 2016, Luleå, Sweden

Proceedings Section/Chapter Finally, any subsidence effect induced instability of completely filled underground mine on topographical surface involving the deep open pit is analyzed. In figure 4, total displacements and failed elements at the end of the production below 1497 can be seen. Orebody settles on the backfilled cut & fill stopes with a deformation of around 270 mm. It is observed that backfilled zones decrease the induced stresses generated by production openings. Total Displacement (mm) 0 mm Ore 300 mm 1497 Level Primary Backfill 1477 Level Figure 4 Global structural state of the mine towards the end of production in cut & fill stopes: total displacements and failed parts Above 1497, 10x25 m sublevel stopes are planned. In the first critical scenario model of the stope layout above 1497, most of the primary stopes are completed and filled. The structural layout for the first critical scenario can be seen in Figure 5. This case simulates the production phase at which maximum load is transferred to the pillar marked in the figure. As can be seen, average factor of safety throughout this pillar is around 2.5. Even for the most critical state, pillar safety is highly satisfactory. Apparently, backfilled stopes decrease the stress concentration on the orebody pillar and increase the factor of safety of this pillar. 1537 Level Ore 2.70 2.60 2.50 FOS vs Distance 2.40 2.30 0 5 10 1517 Level Primary Backfill Investigated Pillar 1497 Level Primary Backfill Figure 5 Structural analysis layout for the critical sequence scenario 1 Another critical state is analyzed by a scenario for the first stoping level above 1497 in which production and filling is completed for all primary and secondary stopes, except the last production stope. Adjacent to this last stope, both sides of the stope consist of primary filling. In Figure 6, plastic model and solution results can be seen. Displacement magnitudes at the stope roof are seen to be in the order of magnitudes Ground Support 2016, Luleå, Sweden 9

Ground Support 2016 proceedings A.F. Surname and J.S.A. Surname less than 100 mm levels. Considering the stope dimensions in orders of tens of metres, total displacements on the stope roof can be regarded insignificant for the overall stability of the stope. Thus, it can be stated that primary and secondary fills around this last stope are effectively performing supporting action as planned. Total Displacement (mm) 0 mm 40 mm Figure 6 Structural analysis layout for the critical sequence scenario 2 At the upper levels, considerable increases in the pillar safety factors are observed and plastic analysis results do not point out any critical deformation states around stopes of the upper levels. Figure 7 represents the surface subsidence after the mining operations are completed and stopes are filled completely. Maximum subsidence is around 140 mm. It is concluded that this much deformation magnitude is not significant for global stability of the mine which has dimensions over 100 metres and depth around 120 m. Total Displacement (mm) 0 mm 50 mm Figure 7 Surface subsidence at the end of backfilling operations and mine life. 10 Ground Support 2016, Luleå, Sweden

Proceedings Section/Chapter Backfill strength is directly related with the cement ratio and aggregate rock strength. In order to provide the global mine stability presented with the numerical models above, aggregate rock particle size should not increase 30 mm and cement mixture should be used as a binder. If the waste rock strength is low, high strength aggregates like limestone should be used. 4.2 Evaluation of Backfill Performance in Practice and Recommendations In mining practice, production started from the bottommost level. After completing a few of stopes, initial backfilling attempts were carried out under control by the company s technical staff. Drill cores were taken from a sample primary backfill operation site for mechanical tests. In coring process, low core recovery was reported. This meant that filling practice was not effective and there were quite large unfilled spaces in filling locations. Drill core samples with 63 mm diameter taken form parts through the waste and cemented mixture were tested in the Rock Mechanics Laboratory. Visual inspections showed that particle size exceeded the recommended -30 mm suggestion. Another issue was identified as the irregular distribution of cement mixture. Improper cementation process was not able to construct strong enough binding among the rock particles and cement. Binding was not sufficiently tight and large voids remained inside. Overall, eight uniaxial compressive strength (UCS) tests were done on cores of this filling texture. Table 4 shows the test results. As can be seen, standard deviation of UCS is high. While maximum UCS is 9.2 MPa minimum UCS is 0.9 MPa. High UCS results correspond to the testing of cores made up of stone boulder parts of the waste. Observations indicate that some core samples are directly taken through the rock and large boulder parts filling the stopes. Thus, test results shown below should not be used to represent large field scale back fill strength characteristics. Instead, core samples taken from backfill with proper particle size distribution and cement mixture should be used to investigate the backfill quality. Aggregates can be obtained from waste rock extracted from the mine. Maximum particle size should be reduced below 30 mm by passing them through crushers. Performance of secondary backfill supplied from the waste rock is analyzed in the modeling work, and it is concluded that if the contact between the backfill upper level and the roof can be sustained with proper backfilling and sufficient yielding of the roof, there is no need to add cement mixture. For primary filling practice, recommendation is to crush the waste rock down to 30 mm with an onsite crusher. In addition, aggregate and cement should be thoroughly mixed before filling into the stope. The mixture should be uniformly distributed, and it should entirely cover the stope space. Although no strength problem for the stone parts of the aggregate is observed, high quality aggregates like limestone available at mine site are suggested to be used in case of low waste rock quality. Table 4 Laboratory test results Sample No Density (kg/m 3 ) Spc1 2.3 Average Density (kg/m 3 ) UCS (MPa) Spc2 2.2 6.3 Spc3 2.3 0.9 Spc4 2.1 3.5 2.2±0.1 Spc5 2.0 4.9 Spc6 2.4 9.2 Spc7 2.3 3.0 Spc8 2.2 1.4 4.1 Average UCS (MPa) 4.2±2.7 Ground Support 2016, Luleå, Sweden 11

Ground Support 2016 proceedings A.F. Surname and J.S.A. Surname 5 Conclusion Backfill design and quality control attempts in Karacat underground iron mine are analyzed. Field investigations and laboratory tests are used to define rock mass characteristics and mechanical properties. Stoping methods are primary choice of any company to lower the production cost. Numerical studies revealed the instability in case of production with a single method, which is sublevel stoping. Lower rock quality and increasing stresses with depth around the pillars and stopes of the footwall levels are identified as the main reasons of the instability. A hybrid underground mining method is proposed and analyzed to ensure stability of local pillar-stope layouts and global mine stability. Below the mine level 1497, production starts by cut & fill method. Completed stopes are filled with the primary type of fill. Above the level 1497, sublevel stoping method with larger stopes is suggested to be applied, considering the economic feasibility of large stopes in mining practice. Observations and mechanical tests were conducted on backfill samples from the field practice. It was observed that aggregate particle size of mine filling applications was much greater than the recommended maximum size of -30 mm. Because of large particle size, cement mixture was concluded to be ineffective in binding the rock particles properly. Low core recoveries reported during core drilling through the filled areas implied that there were large voids inside the supposedly filled stopes. For primary stopes, cemented rock fill is recommended. UCS of this fill is expected to be around 1.5 MPa. To achieve the desired characteristics, mine waste rock is recommended to be crushed below 30 mm particle size and mixed with a cement ratio of 6-8%. For secondary stopes, a rock fill with a UCS of 0.75 MPa is found to be satisfactory. Acknowledgement The authors would like to thank the staff of Özkoyuncu Mining Inc. for their valuable supports. The company is also acknowledged for the permission to use and publish the data. References Alejano, L.R., Rodriguez-Dono, A., Alonso, E., Fernandez-Manin, G., 2009. Ground reaction curves for tunnels excavated in different quality rock masses showing several types of post-failure behavior. Tunneling and Underground Space Technology 24 689 705. Barton, NR., Lien, R., Lunde, J., 1974. Engineering classification of rock masses for the design of tunnel support. Rock Mech. 4, 189 239. Bieniawski, Z.T. 1989. Engineering rock mass classifications. New York: Wiley. Cai M., Kaiser P.K., H., Tasaka Y. and Minami M., 2007. Determination of residual strength parameters of jointed rock masses using the GSI system. International Journal of Rock Mechanics & Mining Sciences 44 (2007) 247 265. Hambley, D. F. (2011). Backfill Mining. In P. Darling, SME Mining Engineering Handbook (pp. 1375-1384). Society for Mining, Metallurgy and Exploration Inc. Hoek, E. and Brown, E.T. 1997. Practical estimates of rock mass strength. International Journal of Rock Mech. & Mining Sci. & Geomechanics Abstracts. 34 (8), 1165-1186. Hoek, E., Carranza-Torres, C. and Corkum, B. 2002. Hoek-Brown criterion 2002 edition. Proc. NARMS-TAC Conference, Toronto, 2002, 1, 267-273. Tiringa, D. (2009). Mining Geology of Karacat Iron Deposit in Kayseri-Yahyali-Karakoy (in Turkish), MSc. Thesis. Ankara: Ankara Üniversitesi Fen Bilimleri Enstitüsü, 139 p. 12 Ground Support 2016, Luleå, Sweden