مجله زمین شناسی مهندسی
سال چهاردهم شماره 1 (بهار 1399)

Drilling has various methods that from different aspects such as crushing mechanism, type of used energy etc., is divided to several types containing hand held drilling, percussive drilling, cable-tool drilling, rotary (or circular) drilling, percussive-rotary drilling and core drilling. Unlike the direct circulation drilling system (DC) in the reverse circulation drilling system (RC), the drilling fluid moves the annulus between borehole wall and the drilling pipe and comes back with the drilled pieces along inside the drilling pipe. The exploratory drilling system of RC by conducting powder samples with high purity and fast drilling rate, is a great help to the velocity and accurate of exploration of ore deposits. Samples produced in this method are in the form of soil and rock powdered and rock fragments of the drilled part, which may be dry or with little moisture. The air flow inside the cycle causes the collected powder sample to be often dry but sometimes is wet due to groundwater or drilling mud. Drilling is one of the most costly mining processes. Therefore, the most important goal in drilling engineering is to reduce costs, and the best possible decision to optimize the cost of drilling is to choose the best possible drilling method. Based on the field data, cost of drilling for each meter of a soft rock (e.g. travertine) by core drilling and direct drilling methods are about 3.3 and 1.2 times of the RC method, respectively. Also the cost of drilling, for each meter of a hard rock (e.g. granite) by core drilling and direct drilling methods are about 2.6 and 1.3 times of the RC method, respectively.

In the present research, reverse circulation drilling (RC) has been compared with other important, common and practical drilling methods, such as direct circulation and core drilling methods in terms of various criteria containing drilling (time) rate, price (cost), type and quality of acquired samples and performance efficiency of drilling. Also, as a field study in this research, deep drilled boreholes with RC and core drilling methods in the gold mine of Khomein-Akhtarchi located in the Markazi province, were investigated and compared from different aspects. At the end, the ability to select the most appropriate drilling method among the variety of methods was studied. The study region is located at 25 km northeast of Khomein city in the Markazi province. This region consists of two exploration areas of Zarmadan-Akhtaran1 with the area of 13.21 square kilometers and Zarmadan-Akhtaran2 with the area of 2.85 square kilometers. Access to the Akhtarchi gold region is possible through the Khomein-Shahabiyeh (Goldsat)-Mahallat road. In the mining region, the Permian rock complexes include dolomite, dolomitic limestone from brown to dark gray, black Irony sandstone and white to milky limestone known as pds, pdl and pl units in the geological maps. In the studied region, several deep boreholes, most of them by RC and some of them by core drilling methods have been drilled. In general, by now in the Akhtarchi gold zone in the Zarmadan-Akhtaran2 area 54 powder boreholes have been drilled through RC method called by RC1 to RC54. Also, there are 25 core drilling boreholes, 18 boreholes called by BH1 to BH18 in the Zarmadan-Akhtaran1 area and 5 boreholes called by BH1 to BH5 in the Zarmadan-Akhtaran2 area. During drilling operations, Permian and Cretaceous rock units have been encountered. The details of drilling via RC method for 4 boreholes with numbers 50, 51, 53 and 54 have been accurately taken. The measured drilling times were obtained from drilling personnel of the mine through the questionnaire which they were weighted mean if needed.

The average drilling time for each meter of rock in boreholes 53 and 54 is 2:12 and 2:54 minutes, respectively. In both cases, the time duration is very short and this feature is one of the advantages of the RC drilling method. The longer average duration of drilling for each meter of rock in the borehole 54 than 53, is due to the depth of the borehole 54 and the hammer problem of the drilling machine during the drilling this borehole. In Table 1, the average duration of drilling operation per meter of rock in the Akhtarchi gold mine is given according to the type of rock (lithology) at definite depth intervals, on the basis of field studies. According to this table data, the duration of the drilling for each meter of rock in the greater depths increases that the reasons for increasing the duration of drilling for each meter of rock in greater depths are the difficulty of drilling due to the increasing length of rig, the reduction of transient energy to the bit, the probability of greater borehole declination, compaction increasing and as a result increasing the strength of rocks and more hydrostatic and lithostatic pressures in the great depths meanwhile at a great depth, the probability of capturing the drilling rig is too high. Also the cost (the time price) of drilling per meter of rock in this mine based on the dip and depth of drilling is about 1300 to 2000 thousand Rials by the RC method, against 2620 to 4250 thousand Rials by the core drilling method. The results of the present research indicate that the RC drilling in comparison with other drilling methods, especially conventional and applied ones in terms of drilling costs and drilling rate (time) is highly desirable while is desirable regarding depth of drilling, the type and quality of the acquired samples and the overall efficiency of drilling performance. Although the core drilling method with the ability to drill very deep boreholes obtaining cores in terms of the type and quality of the acquired samples, as well as the depth of the drilling is the most desirable, but for exploration drilling (especially in the detailed exploration stages), deposits with low-grade and very little mineral indices (like gold mine of Khomein-Akhtarchi), and hence the large sample sizes are needed, employing RC drilling method having comparative advantages is economic.

Regarding the use of RC drilling method in the case study, the gold mine of Khomein-Akhtarchi, it was found that the RC method compared to the core drilling method, in terms of the duration of drilling operations or the speed of advance (the rate of penetration in the rock), drilling costs and efficiency of performance is desirable. Also, according to the type of mineral deposit (gold type), which is low-grade and the indices of the mineral are very low, therefore the large sample sizes are needed, thus, in terms of the type of obtained samples, employing RC drilling method in this case, is accounted a very important advantage related to the DC method (in terms of accuracy) and core drilling method (in terms of cost). The results of this research are useful for all users of drilling operations, including drilling engineers and technicians, engineering geology and geotechnical practitioners, mineral exploration engineers, groundwater aquifers and hydrocarbon reserves (oil and gas) to choose the optimal drilling method under different environmental and economic conditions based on criteria such as the purpose of drilling operations, costs, progress rate, type and quality of the yielded samples and the efficiency of drilling operation. Also, the use of RC drilling method has the advantages over the other drilling methods to be suggested for exploration of low-grade deposits such as gold, silver and copper, especially in the final stages such as detailed and mining exploration.

Studying the effect of slope angle on bearing capacity of foundations on the slope in urban areas is a challenging problem that has been investigated by researchers for years. In general, the analytical approaches for solving this problem can be categorized into limit equilibrium, characteristics and limit analysis methods. In recent years, there have been studies for using the limit analysis within the framework of finite element method for geomaterials. In these studies, the soil mass is not considered as rigid and there is no need to predefine a failure surface for the slope. In the performed research, using the upper bound finite element limit analysis, bearing capacity of strip foundation on slope have been studied. This analytical method enables the use of the advantages of both methods of limit analysis and finite element analysis. In this method, the slip between the two elements is considered. In order to find the critical state of the failure, the rate of power internally dissipated is linearly optimized, by using the interior points method. The advantages of this method are the high convergence rate in comparison with other analytical optimization methods. The effect of different upstream and downstream slopes and foundation depths and also the influence of various mesh discretizations have been evaluated. Finally, the results are compared with those obtained from previous methods available in the literature.

The finite element limit analysis method is based on nodal velocities. Considering the principals of the finite element method and having the nodal velocities, the velocity at each node of the element can be obtained from corresponding shape functions. The rate of power internally dissipated in each element is defined by multiplying the strain rate on stress in each element. In this method, the slip between the two elements and the rate of internal power dissipated at each discontinuity of two adjacent elements is considered. For this purpose, in each node, four new unknowns’ velocities are defined. To remove the stress from the equations, and provide a linear relationship for linear optimization, a linear approximation to the yield function has been used. For this purpose, the Mohr-Coulomb yield criterion is estimated with a polygon in the stress space. Also, using the reduced strength parameter, the effect of the dilation angle is considered. According to the principles of upper bound limit analysis, the value of plastic strain rate is calculated from the flow rule. The velocity field in elements and discontinuities must satisfy the set of constraints imposed by an associated flow rule. In order to have an acceptable kinematics field, the velocity vectors have to satisfy the boundary conditions. These conditions include zero kinematics velocities along the vertical and horizontal boundaries of the geometry as well as negative vertical unit velocities and zero horizontal velocities at points underneath the rigid foundation.

In order to calculate the bearing capacity of foundation, a set of different uniform and non-uniform mesh has been examined. The results obtained from different uniform mesh sizes indicate a certain divergence in the course of analysis. However, the results between the fine and very fine non-uniform mesh are closely related to each other and are converged. The obtained results show that, by increasing the internal friction angle, the bearing capacity has been increased. At high angles of modified friction, the effect of increasing the internal friction angle on the increase in bearing capacity is more in slopes with lower angles. By increasing the downstream foundation depth, the bearing capacity has been increased. This increase is more important in the case of slopes with lower angles. However, the upstream depth variations didn't present a significant effete on bearing capacity. In order to investigate the effect of upstream angle on the bearing capacity, the upstream mesh is also refined similar to the downstream. The obtained results indicate that variations of the upstream angle have a minor effect on the bearing capacity. This is of course true if the upstream slope is fully stable. The results of the proposed method in this study are an upper bound for the results reported by the limit equilibrium and displacement finite element methods. As seen in Figure 1, the suggested method predicts lower bearing capacities compared to rigid block limit analysis method and is indeed a lower bound for the classical limit analysis method. The finite element limit analysis with linear optimization has resulted in more bearing capacity than cone optimization. The bearing capacities, obtained from characteristic lines method depending to the slope angles, in some cases is more and in some cases less than those explored by the proposed method.In this paper, the bearing capacity of foundation located on slope was evaluated by finite element limit analysis method. In this regard, the effects of different downstream and upstream angles of slope and foundation depths and also, the effect of various mesh discretizations on the bearing capacity were studied. It is shown that an increase in the downstream angle causes a decrease in the bearing capacity and an increase in the downstream foundation depth leads to an increase in the bearing capacity.  However, the upstream angle and upstream foundation depth were not much effective on the bearing capacity.

Because of the diversity in petrography, peridotites have variable physical and mechanical properties. For this reason, knowledge of resistance properties and their deformation will help with the prediction of engineering behavior of these rocks. Due to the large spread of igneous rocks, especially peridotite, in Zagros, northeastern and central Iran, special attention has been paid to their petrographic, physical and mechanical characteristics. The construction of the structure within or on the peridotites and the choice for the purpose of the stone borrow depends on the recognition of its engineering geology characteristics. In this paper, in addition to the field and laboratory study, the geological characteristics of peridotite engineering has been investigated.

In order to study the geological characteristics of the peridotites of Harsin region, 15 suitable blocks were selected and transferred to the laboratory. Accordingly, from collected rock samples, 150 cylindrical cores of diameter 54 mm were prepared and physical and mechanical tests were performed according to (ISRM, 2007) and (ASTM, 2001) guidelines. In this research, after sampling of the study area and preparing the core for the lithological characteristics of the samples by providing thin sections of them with polarizing microscopy was studied.

By considering the results of laboratory tests and analysis from Harsin peridotites in Kermanshah province, we can acclaim that with increasing the percentage of minerals in olivine and pyroxene in rock, the strength was decreased and the levels weaknesses, which is due to the weak structure of the mineral-olivine and pyroxene. According to the physical properties test and Anon classification, the porosity percentage in porosity percentage is low and as a result the amount of water absorption index is low. Based on the Gamble classification, all peridotites are very resistant to durability and based on the Franklin and Chandra classification, all samples are extremely resistant. The results of this study showed that the single axial compressive strength, elasticity modulus, point load index and tensile strength were decreased with an increase in humidity content of peridotite samples. This is due to the fact that with the increase of humidity pore pressure of water increases. According to the Anon classification, the peridotites are very high in terms of the length of the longitudinal passage through the rock. The highest compliance between the Brazilian Tensile strength test (BTS) and Schmidt hammer (SHV) was achieved in the dry condition and the determination coefficient (R2) equals to 0.95 was obtained. Also there is an acceptable relation between the Brazilian Tensile Strength Test (BTS) and the dry volume unit weight (γd) with the determination coefficient (R2) of 0.93. In addition, there is an admissible relationship between durability test and single-axial compressive strength, with a coefficient determination (R2) of 0.94. Regarding the obtained regressions in this study, the physical and mechanical properties show good agreement and most of the equations have an acceptable coefficient determination.

Geofoames are used as a light weight fill material in those places which soil borrows is not cost effective for engineering or economic purposes. In general, geofoames are highly capable of improving some of geotechnical properties of soils such as inflation creation, reduction of density, and etc., due to their light weight, no change of volume against water, low permeability, and relatively proper strength. Using mixture of geofoam beads and soil has been recently taken into consideration by researchers. The mixture causes tangible reduction of soil density and severe drop of active pressure of retaining walls. Also, using the mixture in seismic zones is of special importance. In the paper, effect of mixing geofoam (4 different percent) and three types of poorly graded sandy soils have been dealt with. The research innovation has been compared to previous ones is using poorly graded sandy soil, separating geofoam beads based on their diameter, and reviewing the effect of adding various percentage of geofoam on improvement of poorly graded sandy soil’s properties.

Tests have been performed in direct shear box (10 cm x10 cm) under three stress levels of 50, 100, and 150kPa. First type of soil has been Firoozkooh sand (#161) with specific gravity of 2.65, as uniformly graded sand (SP). Second type of soil has been mixture of uniformly graded sand and 10% silt (SM-SP); and, third type of soil has been mixture of Firoozkooh sand and 20% silt (SM). The three above types of soils have been named as soil 1, soil 2, and soil 3, respectively. Geofoam beads have been all fine grained, passing through sieve No. 10; and, their added weighted values have been 0, 0.2, 0.4, and 0.6% of weighted percentage of soil. All of tests have been performed with optimum moisture content of geofoam and soil mixture. Due to diversity of soil types and ratio of geofoam-soil mixtures, soil compaction test has been performed on each direct shear test’s sample to specify optimum moisture content of various types of mixtures; because there have been various types of soils used, and also various ratios of soil and geofoam mixtures.

According to the results, using geofoam beads leads to considerable reduction of soil density. Decrease made in density will be more tangible when higher percentages of geofoam are added to the soil. Also, as far as geofoam absorbs water, optimum level of moisture will be increased through increase of geofoam percentage in soil-geofoam mixture.Since geofoam beads are less rigid compared to grains of sand, sand and geofoam interlocking and friction level is lower than sand interlocked to sand; and shear strength has been decreased through increase of geofoam percentage in soil. The point to be remembered is that, reduction level of shear strength in soils containing various percentages of geofoam is not so tangible compared to the soil itself. In its worst case, the reduction would be about 12%.Adding geofoam beads to all of the three types of soil has led to their increase of apparent cohesion. Moreover, through increase of mixture percentages, more increase has been made in apparent cohesion of mixture. The results are indicative of significant effect of mixing geofoam and soil 1 in increase of soil cohesion up to 9 times. The cohesion increase has been about 4 and 2 times for soils type 2 and 3 respectively. So, it could be concluded that the lower the soil cohesion, the higher would be effect on cohesion increase of soil, through increase of geofoam percentage.In figure 1, chart of internal friction angle is shown based on mixture percentage of geofoam for those types of soils being tested. Considering decrease of internal friction angle through increase of geofoam percentage, the important point is slope drop observed when geofoam percentage added has been 0.4%. Therefore, reduction speed of internal friction angle has become slower, after this level. Considering the figure, internal friction angles of soils type 1, 2, and 3 have shown respectively 15, 16, and 18% of reduction, through highest percentage of geofoam added (0.6%).Figure 1- Internal friction angle based on geofoam percentage mixed with different soils Comparing the results from present and previous researches, it could be concluded that adding higher percentages of geofoam results in cohesion increase of sandy soils; however, the increase level is different for various types of soils. The lower the initial cohesion of sandy soils and the more uniform their gradation, the more the effect of adding geofoam on increase of cohesion coefficient of soil. Also, downward trend of internal friction angle for well graded and poorly grades sandy soils is almost similar. Using the results from present research and considering acceptable level of reduction made in internal friction angle of the soil mixed with geofoam against cohesion increase and reduction of soil density; mixture of geofoam beads and soil could be used in construction of embankments, retaining walls and other earth structures, appropriately.

The increasing rate of construction activities in urban areas is accompanied by excavation in the vicinity of existing structures and urban utilities. This issue has highlighted the importance of constructing protecting structures in order to control displacements and prevent damage to structures and their neighboring area. Among the important widely used wall stabilization techniques, one can name nailing and grouted anchors. However, these methods suffer some drawbacks such as annoying noise and vibration during the drilling, implementation difficulties below the water table, grouting problem, installation of strands and bars in the borehole in porous and collapse soils, and long curing time for the grout of post-tension anchors. Since the helical anchor method lacks many of the mentioned problems, it is now widely used in many applications.In the present work, a laboratory model of helical anchor stabilized wall is presented and evaluated. For this purpose, four types of anchors at 20° back slope are designed in a sandy soil and the effect of helix configuration (in term of its diameter and number of blades) is investigated. Considering the laboratory scale of the designed model, the results obtained using helical anchor were compared with numerical results of soil nailing wall by applying the particle image velocimetry (PIV) analyses.

The test box designed in this work is made of a metal plate with a thickness, length, width, and depth of 1.5 mm, 100 cm, 60 cm, and 30 cm, respectively, and a Plexiglas in its opposing side with a thickness of 50 mm. The soil used in the experiments was the dry sand of Soufian region in east Azerbaijan province of Iran. The soil is classified as SP according to USCS classification. The helical anchors were fabricated by welding the helical pitches to a metal shaft. The end part of the shafts is screw threaded such that to fasten a bolt to them. To start the experiment, the empty box was completely cleaned using the detergents to remove any pollution or soil on the Plexiglas and metal surface. Afterward, the sandy soil was poured on the wall floor and the facing was placed inside the box vertically. Again, the sandy soil was poured from both sides of the facing up to the installation height of the helices. Helices were installed in the assigned holes and their angle was adjusted through the pre-fabricated stencils. The soil height was increased up to the next row assigned for helices installation. These steps were repeated until reach the wall crest. After preparation of the physical model, its behavior during the preparation must be modeled. We first filled both sides of the model and then modeled the stability behavior of the helical anchor wall through excavating its facing opposed side. Overall, the wall was built through eight excavation steps.

The maximum displacement is related to the anchor type 1, which does not have enough bearing capacity under surcharge conditions. By changing the anchor type and increasing the number of helices, shear strains and their expansion in the wall back decline. The decrease in displacement rate by changing the anchor from type 1 to type 2 is 18%, which is due to the low bearing capacity of type 2 anchor compared to the type 1 anchor. Increasing the number of pitches from one to two (changing the type 1 anchor to type 3 anchor) showed a considerable decrease (i.e., 43%) in displacement rate. Increasing the number of pitches from 1 to 3 (changing the anchor from type 1 to type 3) resulted in a 62% decrease in wall crest displacement. This displacement decrease rate seems to decline with an increase in the number of helixes.The displacement rate for all four anchors is almost similar in two excavation steps, which probably is because of the need for displacement for activation of the anchors. One strategy to deal this issue in the sensitive projects and control the displacement is to apply post-tension helical anchors. Then, in stages 4 to 6, the displacement was almost constant due to four main reasons including wall rigidity, the presence of reinforcements, formation of pre-step displacement-induced tension force, and enough capacity of anchors to face with more displacement. In stages 6 to 8, type 1 and 2 anchors showed growing displacements due to the reduction and ending the wall rigidity and lower bearing capacity. In type 3 and 4 anchors, the maximum displacement was related to 4 initial stages. In type 1 and 2 anchors, which have two helical plates, almost a similar behavior was observed until stage 6 of excavation, but eventually type 3 anchors showed better performance because of higher bearing capacity to overall displacement.

In the present study, a physical model was designed to investigate the effect of helical anchors’ geometry on displacement rate of helical anchor wall and compare it with a nail wall. Overall, comparing the results obtained by conducting these experiments on a helical anchor stabilized wall and a nail wall revealed that:- Wall crest displacement is affected by the diameter and number of helices and decreases by an increase in bearing capacity. The increase in the number of pitches from one to two (single-pitch to double-pitch anchor) has a higher effect on displacement control compared to the case of changing the double-pitch to triple-pitch anchor. So, it can be stated that a further increase in the number of anchor pitches results in a declined performance of the anchors. All anchors need a slight displacement for activation. This issue cannot be resolved by changing the type of helical anchors. Hence, when the displacement required for activation of the anchors exceeds the allowable wall crest displacement, use of post-tensioned helical anchors is recommended.A comparison between nailing and helical anchor results revealed that the relative density of the wall stabilized with the helical anchor is less than that of the nail wall; and wall crest displacement in the helical anchor wall was very lower than that of nail wall. Thus, the helical anchor wall stabilization is preferred when other economic and technical requirements are met.

Rock mass deformation modulus is one of  the major parameters has to be considered in the design phase of arch dams. Due to filling and discharging of reservoir and corresponding loading and unloading on the dam abutments, irreversible deformation takes place within the rock mass and consequently, increases the potential of creating a separation between dam body and abutments. Therefore, the rock mass modulus must be more than an alowable value in order to prevent arch dam failure. Regarding small core samples and lack of joints and other similar discontinuities in samples, the determined modulus through performing laboratory tests is higher than those obtained through in-situ tests. The available technique to estimate the rock mass deformation modulus is divided into two classes as direct and indirect methods. In direct methods, the rock mass deformation modulus is measured via performing in-situ tests such as plate loading test while it is estimated through empirical equations using rock mass classification and laboratory test results in indirect methods. These equations are developed based on regression analysis between the rock mass modulus calculated via in-situ tests, the rock mass classification and laboratory test results. Although application of these equations is simple and cost-effective, the results are doubtful and cannot be used in the design phase of arch dam due to the heterogeneous nature of rock mass and rock type variability. The numbers of micro-cracks which are developed after gallery excavation using drilling and blasting technique are more close to the loading plate. Thus, calculated modulus in these points is lower than reality. The displacement in the points far from loading plate was near to zero while the transmitted load which is calculated applying ASTM D4394 standard is more than reality in small galleries. Consequently, the calculated modulus was extremely larger than real values and sometimes even more than intact value. The empirical equations are site dependent and they are just applicable in sites with similar geotechnical condition. It is obvious that in-situ tests, such as plate loading, are the appropriate method in order to determine the modulus of deformation, however, due to some simplification in the data processing such as semi-infinite boundary condition, the application of numerical simulation as a data processing tool is more appropriate. In this research, the Beheshtabad dam was introduced and the geology characteristics of dam site were investigated. Applying direct and indirect methods, the rock mass modulus of dam abutments is calculated.

The dam site is placed approximately at a distance of 2.7 km from the intersection of Koohrange and Beheshtabad river. In accordance with geological studies, the rocks in the site could be categorized in four units combined of Dolomite, Dolomitic Limestone, Limestone, Marl and Marly Limestone. Applying empirical equation the rock mass modulus of dam abutments is evaluated based on the laboratory test results and rock mass engineering classification systems. In addition, ASTM D4394 is applied to investigate the results of ten plate loading tests which are executed in the right and left abutments. To interpret the plate loading test results in the right abutment, a three-dimensional Fast Lagrange Analysis of Continuum (FLAC3D) model is developed.

To process the numerical simulation results, back analysis as a data processing tool is used. In this approach, the input parameters of numerical model will be changed in the way that the measured quantities by extensometers at the monitoring points are almost equal with the computed ones via numerical model at the corresponding points. Based on the sensitivity analysis carried out on the Mohr-Coulomb failure criterion parameters, the friction coefficient and cohesion variation do not affect the displacements calculated via numerical simulation as the more portion of gallery displacements are elastic. The error function is minimum when the rock mass modulus is 12 GPa and the horizontal to vertical stress ratio (K0) is equal to 0.5. The evaluated rock mass modulus based on the numerical simulation is two times lower than corresponding one evaluated applying empirical equation as a result of empirical equation uncertainty. Consideration of stress decrement under loading plate shows lower level of stress decrement under loading plate in ASTM D4394 compared to numerical simulation. This is why, the rock mass modulus, calculated based on ASTM D4394, increases dramatically by getting distance from the loading plate. 

The empirical methods estimating the modulus of deformation based on rock mass classification systems tend to evaluate large value of modulus especially for the weak massive rocks.As a result of galleries dimensions and semi-infinite boundary condition assumed in ASTM D4394, the calculated rock mass modulus increases dramatically by getting distance from loading plate. Therefore, the numerical simulation was applied to process the plate loading test results. A new normalized error function was developed based on measured displacements and the rock mass modulus in the right abutment was determined 12 GPa which is very lower than the calculated value using ASTM D 4394. Also, as a result of numerical simulation, the rock mass is uniform. The stress increment perpendicular to the loading plate was calculated applying numerical simulation which is 0-90 percent lower than those suggested by ASTM D 4394.