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By A. G. Doll

"This paper expands the SAG mill survey database published by the Author at Procemin 2013. The new database includes 49 published SAG/AG mill surveys and includes mills from all over the world. The paper will compare the survey database to SAG mill models by Austin, Morrell and Hogg & Fuerstenau with the purpose of demonstrating the degree of fit of the models to surveys and validating any empirical “fitting factors” used in the models.SAG mill power draw models are used in mill design and grinding circuit modelling to predict how much power will be consumed by a particular mill geometry and operating configuration. The importance of conducting surveys suited to model calibration will be highlighted because the review of literature shows that survey information important to modelling is often missing or omitted. A recommendation for the collection of data during a mill survey is presented, along with some assumptions used by the Author in the absence of certain data. The benefit to mining companies of publishing their survey data, and thereby allowing modellers to improve their model calibration, is discussed.The comparison of models requires a discussion of the measurement of power in a mill drive system as the models use slightly different position in the electrical network for ""where power is measured"". The benchmarking of models against plant operations requires a similar discussion of power measurement."

Jan 1, 2016

By A. Atrafi, L. Barroilhet

The top size of particles in the flotation of potash ores, which can be as large as 4 mm, exceeds by far the top size in other flotation plants. It is then understandable that while long chain amines are used in the flotation of fine fractions (below 0.8 mm) of potash ores the coarse particle flotation also requires an extender oil. While new and very promising fluidized-bed flotation machines (e.g. HydroFloat) constitute a significant development in the flotation of coarse particles, in the area of extender oils, as reviewed in 2007 by Strathdee, Gotts and McEachern, there have been no new developments to replace the industry standard asphaltene-based carcinogenic extenders. This situation first of all results from the experimental problems that arise when water-soluble minerals like sylvite are to be tested in saturated brine. Both oils and longchain amines are water/brine insoluble. The oil with the amine dissolved in it can however be emulsified and then used as an extender oil. Such oil droplets must wet KCl particles and the emulsion must be stable to avoid handling problems. In this paper we test the use of the phase inversion technique as a method of studying wettability of KCl particles by the oil in saturated brine, and the effect of amine dissolved in this oil on the wettability. The results show that this relatively simple technique can be used to characterize the wettability of fine KCl particles by the tested oil and to determine the effect of added amine on it, and so it can be useful in selecting proper extender oils.

Jan 1, 2016

By Balmer Neilly

In these days of arrested prosperity, when patience is tried and the future but dimly outlined, Canada's gold mining industry stands out as a beacon light shedding its rays of hope and confidence and demonstrating that there is still one product unaffected by Customs barriers and eagerly sought by all the people of the world. Some other forms of industry have hesitated and side-stepped. Some have cried for help, and in spite of public assistance, generously extended, they are still unable to demonstrate definitely their ability to resume the onward march. But gold mining has never hesitated. As an industry, it has bad an unique experience. When prices are high and conditions are generally referred to as prosperous, this industry is adversely affected. The cost of production goes up; profits are diminished; public sympathy is lacking; and, thrown upon its own resources, the industry tightens up its belt and throws its whole energy into an effort to improve operating conditions, and offset high prices by increased efficiency. When, in due course, depression follows inflation and prices ebb, the gold miner, through necessity, finds himself physically and mentally prepared to meet the situation.

Jan 1, 1932

By H. Clare Horwood

"Fractures and fracture zones containing veins and stringers of auriferous quartz in a quartz porphyry dyke are the principal structural features at the Howey and Hasaga gold mines. The 'break' containing the dyke was produced by a force from the northwest, the fractures in the dyke by compressive stress from the west.Both mines are in the north-central part of Heyson township, in the Red Lake area. They lie close to the southeastern side of the Dome granite stock. The Howey mine exhausted its ore reserves and ceased operations in November 1941. The Hasaga mine is producing at the present time (June, 1946).Mineralization was first discovered in the dyke in 1925 by the Howey brothers. Lorne Howey's discovery, which became the Howey mine, came into production in April, 1930; Ray Howey's find, which is now the Hasaga mine, started production in November, 1938.The total production from the Howey mine was 421,324 oz. gold and 144,253 oz. silver from 5,158,376 tons of ore hoisted ( 4,319,331 tons milled) during its 11 years of operation.The total production from the Hasaga mine from November, 1938, to December 31, 1945, was 124,434 oz. gold and 51,009 oz. silver from a total of 901,098 tons hoisted (721,361 tons milled).GENERAL GEOLOGYThe porphyry dyke containing the orebodies at the Howey and Hasaga mines lies a short distance southeast of the contact of a granite-granodiorite mass known as the Dome stock. The dyke fills a fracture trending N.65°E. in a tongue of Keewatin volcanic breccia that extends in a rather irregular manner into an elongate intrusive stock of dioritic rock known as the Howey diorite. The accompanying table presents the geological succession."

Jan 1, 1954

By J. -C. Parisien-La Salle, D. Blais, G. P. Demopoulos

"In this work the dissolution, precipitation and adsorption behaviour of gold in acidic iron sulphate media containing small amounts of chloride ions is investigated via batch laboratory tests and OLI thermodynamic calculations over the temperature range 25-250°C. Chloride ions may be present in industrial pressure oxidation/pressure leaching (POX/PL) operations due to elevated water salinity levels or due to intentional addition as is the case of the CESL Copper Pressure Leaching Process. Thermodynamic calculations show that AuCl4 - may form depending on the free Cl-/Au ratio and temperature, its stability increasing with increasing Cl- concentration and decreasing with increasing temperature. On the other hand during POX/PL, we know that massive precipitation of iron in various hydrolysis products and compounds occurs. Hence of interest is the question how AuCl4 - that may form during leaching interacts with the iron precipitation products. Gold(III) chloride was found to co-precipitate with iron(III) (oxy) hydroxides and arsenates forming in the autoclave but not with natrojarosite. Upon further investigation it was discovered that activated carbon added in the autoclave (as done in CIL) acts as preferential sink for the mobilized gold, hence opening the potential for the development of a direct cyanide-free gold recovery process in acidic sulphate leach reactors.INTRODUCTIONThe recovery of gold from ores and concentrates has evolved over the years with processing routes becoming increasingly complex. Given the high value of gold, any reduction in losses during recovery operations constitutes an important economic benefit. One such loss that needs to be addressed is the potential solubilisation of gold during pressure oxidation (POX) of refractory sulphidic gold feedstocks. For example, Agnico Eagle Mines reported a loss in gold recovery due to the formation of a gold chloride complex (Garofalo, 2009). Gold is known to form complexes in solution, such as gold(III) chloride, if an excess of chloride ions are present. The fate of such soluble gold chloride during pressure oxidation and the subsequent neutralization of the acidic discharge slurry will determine if gold is lost or recovered in the downstream cyanidation operation. Thus it is of interest to know under what conditions gold chloride may stay soluble or co-precipitate along primary (formed in the autoclave) or secondary (formed during neutralization) precipitates. Furthermore, the mode of co-precipitated gold, such as adsorption on iron compounds, like goethite and hematite (Machesky et al., 1990; Nachayev, 1984); substitution in one of the precipitated phases; or finally chemical reduction to metallic state would have consequences in terms of gold recovery."

Jan 1, 2012

By M. K. Walter

"The underground operations of Kidd Creek Mines uses an average of 450 U.S. gpm water during regular operations, and discharges an average of 800 U.S. gpm through consecutive staged pumping lifts of 1200f t.Water is used as a flushing medium for the drilling operations and as a dust suppressant for drawpoint mucking. The water retains an average of 5% suspended solids by weight which is removed underground before being pumped to the surface. This paper outlines the treatment of the mine water for the removal of suspended solids which are chemically treated to settle out as a sludge product of 60% solids by weight. This process reduces the sludge handling operation by 50%. The sludge is systematically disposed into the ore passes.IntroductionThe Kidd Creek minesite, located 27 km north of Timmins, Ontario, produces 4.5 million tonne s of copper, lead, zinc and silver ores annually from its No. I and No.2 underground mining operations. Both mines use a sublevel blasthole stoping method of extraction.Water from the underground operation passes through a series of surface settling ponds, where it is treated to remove heavy metal ions from solution. A blend of the pond effluent with fresh water from Kidd Creek, supply the underground with water at an average rate of 450 U.S. gpm. Water flows from the production zones through coarse sumps and into a main pumping station. There are five main pumping stations underground, three in No. I Mine and two in the No.2 Mine.The No. 2 Mine pumping stations are located on the 4800 level and the 3600 level. All water below the 3600 level to the shaft bottom is controlled from the 4800 pumping station. The water from the 4800 pumping section is combined with water from the remaining levels of the No.2 Mine at the 3600 pumping station. Water from the 3600 pumping station is pumped to the 2400 pumping station in No. I Mine.The No. I Mine pumping stations are located on 2800, 2400, and 1200 levels. All water below the 2400 level to the shaft bottom is controlled through the 2800 pumping station. This water is pumped to the 2400 pumping station. The 2400 pumping station also controls the water below 1600level to 2400 level."

Jan 1, 1988

As a bubble travels through the pulp region of a flotation cell, it will collect the most hydrophobic particles of an intermediate particle size range, this being the primary source of true flotation selectivity. Despite its role on bubble-particle attachment, however, particle hydrophobicity is still missing a mathematical definition that can be used for plant simulation. In this paper, particle hydrophobicity is defined in terms of the induction time necessary for attachment and then linked to the mass of collector adsorbed per particle. This allows particle hydrophobicity to be written as a function of density of collector adsorption as well as particle size and mineralogy. The model is calibrated from the experimental measurement of the flotation rate constant. Liberation data from an industrial flotation rougher is used for model validation. It is demonstrated that the model can describe a broad experimental dataset by means of a minimal number of parameters. Main industrial application is in simulating the effect of the chemical environment upon plant metallurgical performance.

Jan 1, 2006

By John Russell

"Faster, safer, grinding mill relining offers the obvious attractions of reduced mill relining times with the knock-on incentive to reduce all other maintenance down-time, opening the possibility for increased annual plant utilisation.Leading edge liner handler technologies by RME, focused toward even shorter relines for very large SAG and AG mills, have recently split into two quite distinct paths:•,?Single SAG/AG Mill Relining Machines (MRM’s) of very large capacity, such as Newcrest Cadia’s 40’ SAG MRM of 6000 kg capacity and P.T. Freeport’s 38’/34’ SAG MRM of 7250 kg capacity;•,?Twin-machine SAG/AG Mill Relining Systems of moderate to large capacity such as BHP Billiton’s Escondida Laguna Seca 38’ SAG MRM’s, each featuring 4500 kg capacity.RME, its customers noted above and their liner suppliers are providing leadership in this important annual utilisation advance.Once the 24 hour mill shell reline becomes reality, liners designed for maximum profit, not maximum life, albeit exchanged more frequently, become possible too.INTRODUCTIONAt most mineral grinding plants, mill relining activities dictate each maintenance shutdown’s critical path. When annualized, mill relining remains the largest, single influence in defining plant availability.Because of restricted access, both inside and outside the mill, increased application of manual labour has no effect on reducing relining downtime. The only answer for RME has been to develop and apply new relining technologies now known as RME’s Mill Relining System.RME’s primary focus for over twenty years has been the reduction of grinding mill relining times while simultaneously providing a marked improvement in relining safety. Our original (1990) target was to halve traditional (pre-RME) reline times and to eliminate accidents and injury by mechanizing all physical work."

Jan 1, 2008

By Kish. L.

"The results of a chemical and mineralogical study of eighteen opaque-oxide occurrences are discussed and the detailed chemical, metallurgical and mining feasibility study of a vanadium-rich magnetite deposit is summarized. A plot of the mol-percentages in the system FeO- Fe2O 3-TiO2 indicates that the deposits fall into two distinct groups: some contain hematite with ilmenite, and some others have magnetite with ilmenite (titaniferous magnetites). The plot also suggests that the two types of deposits were crystallized from liquids of different compositions. Each group has its distinct structural features. Vanadium is present in all of the studied specimens, however concentrations of economic importance have so far been found only in the titaniferous magnetite deposits. The vanadium content of the oxide concentrations of the Dore Lake Complex is the largest among the studied occurrences. Analyses of the concentrates of the drill-core sections show a compositional variation across the orebody; a continuous increase in the tenor of Ti02 and a decrease in the vanadium content upward in the layered sequence is found. Concentration tests on several tons of samples indicate that about 80 per cent of the vanadium is recovered in the magnetite concentrate."

Jan 1, 1972

By Helen Shang, Jianjun He

In nickel and copper production, sulfur dioxide (SO2) is generated during mineral smelting. Emission of sulfur dioxide into the atmosphere causes serious health and environment issues. Sulfuric acid plants, which convert the harmful SO2 gas to commercially useful sulfuric acid product, ensures safe and environmentally sound smelter operations. The sulfuric acid plant is mainly composed of a SO2 to SO3 converter, a series of heat exchangers and SO3 absorption towers. The SO2 converter is crucial for minimizing SO2 emission in a smelter. In this paper, steady state and dynamic modelling are carried out for the converter using mole and energy conservations. Simulations are performed to investigate the effect of feed conditions on SO2 conversion. The developed model represents an important component for modelling the acid plant in a smelter and is expected to provide a fundamental tool for the future optimization and control development.

Jan 1, 2014

By L. S. Trenholme

"Belleterre Quebec Mines, Limited, is a subsidiary of Mcintyre Porcupine Mines, Limited. The property is in Guillet township, Quebec, 45 miles due east of Haileybury, Ontario. It is accessible by 72 miles of gravelled road from New Liskeard, Ontario (Fig. 1).Gold production at the property commenced in 1936 and to March 31st, 1945, had exceeded $9,000,000. For the past 3 years the milling rate has averaged 325 tons a day and the gold recovery about $14 a ton.GENERAL GEOLOGYThe rocks of the district consist of Keewatin-like volcanics conformably overlain by Timiskaming-like sediments. Together they form the remnant of an easterly plunging anticline engulfed by granite (Fig. 2). The pregranite intrusives consist of diorite, gabbro, and ultrabasic rocks, occurring as dykes, sills, and irregular stocks. Dykes of porphyry and lamprophyre presumably related to the batholithic intrusive are common in this area. The veins of quartz are also probably derived from the batholithic intrusives. A few dykes of Keweenawan-type diabase have been observed."

Jan 1, 1954

By Michael L. Free

The hydrometallurgy curriculum for undergraduate students at the University of Utah has evolved to contain a variety of tools to help students understand and apply hydrometallurgy in a context that has industrial relevance. In order for students to perform as engineers in an industrial hydrometallurgy setting, they need to understand thermodynamic and kinetic fundamentals and be able to apply these fundamentals to extraction, concentration, and recovery applications. Students must also understand related environmental, economic, and assessment issues in hydrometallurgy and be able to work with others. In order to meet these needs, the hydrometallurgy curriculum includes a variety of relevant subject information, on-line resources, laboratory experiences, and team work opportunities to help prepare students for careers involving hydrometallurgy. The development and use of these resources will be discussed in this paper.

Jan 1, 2014

By Y. Chunhua, L. Chunsheng, W. Sheng

"Countercurrent solvent extraction is the dominant industrial rare earth separation method. For the rare earth separation, with low separation factors and multiple components, designing and optimization plays an important role in reaching product quality and reducing separation plant costs. Based on the theory of countercurrent extraction, cost-efficient methods, such as third-outlet separation and hyperlinked extraction processing, have been introduced following computer simulation. The consumption and fixed investment of industrial practice are compared, by which advantages of the optimized rare earth separation process are indicated. INTRODUCTIONThe rare earths (RE) have outstanding optical, magnetic and electric properties in advanced material applications and have therefore received much international attention in recent years. Liquid-liquid extraction is the major industrial RE separation method and is widely used in the industry. Supported by the government and with collaboration from the institutions and enterprises, China has made great progress in RE separation, and has become the dominant supplier of separated RE products. Complicated by multiple elements and low separation factors, careful design of RE separation processes is vital for the quality and cost of production. The theory of separation is the key to basic of flowsheet design and industrial applications.BASIC SEPARATION PROCESS AND THEORYGeneral ProcessAcidic extractants such as P507, P204 and naphthenic acid, are the most popular extractants in modern RE separation industry. In the general separating process, the extractant is saponified with a base, then ion-exchanged with RE solutions, thereby to provide stable and high RE loading in the extractant to fulfill the requirement of production quantity and quality. The RE-loaded extractant is then countercurrently extracted and scrubbed with RE solutions and the RE element thereby is purified. After the countercurrent extraction process, the RE-loaded extractant is stripped with acid to obtain an aqueous RE product, and prepare the extractant for recycling (Xu, 1995)."

Jan 1, 2012

By P. M. STILES

Acquired in 1957, the limestone quarry of Ideal Basic Industries on Texada Island north of Vancouver has become a major supplier of high purity limestone and construction stone. Crushed limestone products are shipped by barge to various destinations along the B.C. and northern U.S. coasts.

Jan 1, 1984

By G. R. MANSON

The Brisco Barite Mine lies 70 km south-southeast of Golden, British Columbia. Barile mineralization occurs in a northstriking brecciation zone within Ordovician-Silurian Beaver/ oot dolomites and quartzites. Brecciation was caused by faulting related to the Mount Forster-Steamboat Fault. The main barite vein occupies a fault zone and is up to 12 m wide and over 250 m long. It is light grey to white, fine grained and compact; BaSO4 content is 9/. 70 per cent. Production started in 1945 and has continued to the present. Barile originally was removed from an open pit; later, an adit, begun in 1957, reached more than 240 m along the vein, 15 m below the pit floor. Today ore is hoisted from a decline resting some 37 m below the pit floor.

Jan 1, 1984

By A. H. Winckers

"The Sullivan Mine and Concentrator are owned and operated by Cominco Ltd. and are located in the Rocky Mountain Trench, in Kimberley, British Columbia.The Sullivan Concentrator commenced operation in 1923 at approximately 2,500 tons per day. The concentrator was enlarged in several stages and by 1945 was treating 8,500 tons of ore per day. In 1949, the capacity of the concentrator was expanded to 10,000 tons per day with the completion of the sink and float plant, the installation of the rod mill, an underground crushing plant and a 4-mile underground and surface electric railway from the underground crushing plant to the concentrator. Throughput now averages 7,000 tons per operating day, or 2,400,000 tons per year.The concentrator operates 7 days per week, 24 hours per day while the underground crushing station operates 5 days per week, 24 hours per day. The ore is essentially a replacement deposit in argillaceous quartzite of Precambrian age. The ore bodies are massive deposits consisting of fine grained mixtures of sulphides, sometimes interbanded with the country rock and varying widely in their content of lead, zinc, iron and silicates. The principal economic minerals are galena and marmatite (7 ZnS:FeS). Iron is mainly present as pyrrhotite, and to a much lesser extent as pyrite. The intimate association of these sulphides necessitated the development of a successful differential flotation process before the ore could be beneficiated profitably."

Jan 1, 1981

By A. F. Matheson

"Bankfield Consolidated Mines, Limited, and Tombill Gold Mines, Limited, in the Little Long Lac area, Ontario, mined an orebody that was common to both. The discovery' of this orebody was made by surface prospecting in May, 1932, on Bankfield ground, following an original discovery made in the same manner on the same property on a small reef in Magnet lake in October, 1931. The production from the two mines between 1937 and 1942 was approximately 135,000 oz. gold from about 420,000 tons of ore, an average grade of 0.32 oz. a ton. Production was nearly equally divided between the two properties. Both mines are now closed, the ore exhausted.""The ore is mainly sheared and fractured greywacke and porphyry impregnated by silica and small amounts of metallic minerals. To a lesser extent there are definite quartz veins in the ore zone""(l). The metallic minerals present were principally pyrite, arsenopyrite, pyrrhotite, and native gold .As at other mines in the Little Long Lac area, an elongated, broad, deep trough of folded, steeply dipping volcanic and sedimentary rocks, striking roughly east-west and surrounded by younger granite, is host to the orebodies of the Bankfield and Tombill. In the immediate area of the two mines iron formation and conglomerate bands occur nearby both north and south of an intrusion of porphyry-Cliorite, suggesting repetition of beds and folding. Although considerable drag folding is in evidence, folding on the above scale has neither been proved nor disproved."

Jan 1, 1954

By Jos. Kalbhenn

PART of the plan of re-allocation evolved by the engineers of the Dominion Steel and Coal Corporation in 1932 to retard the rapid advance of the submarine workings seawards was the opening of Dominion No. 20 Colliery, in the Harbour seam, by a pair of cross-measure tunnels from the workings of an older mine working the Phelan seam, which lies 430 feet below the Harbour seam. The tunnels, 8 ft. by 13 ft. in size, were started away from the old workings of Dominion No. 2 at a point 1.5 miles beyond the shore line and were driven on a rising grade of 1 per cent for a distance of 5,335 feet, tapping the Harbour seam at a point 2 ¼ miles off-shore. For transportation and ventilation, shafts and roadways already existing in the Phelan seam were made use of during the stages of development and up to the present time. One of the two tunnels served as a main haulage road and an intake airway, and the other as a return airway and travelling road. However, as the workings of this new colliery expanded, the need for additional airways arose as was anticipated when the plans for the development were formulated. To meet this need, work was started on October 1st, 1942, to establish a new opening between Nos. 2 and 20 Collieries in the form of a vertical shaft to provide a passageway for the ventilating air current of some 250,000 cubic feet per minute on its return to the upcast shafts in No. 2 Colliery, thus making both tunnels available for intake airways. In the interest of safety, speed, and economy, the shaft was raised from the Phelan to the Harbour seam. It is 20 ft. by 12 ft. in size and, with the exception of three timbered compartments across the narrow end, is untimbered. One of the three timbered compartments housed the ventilating tubing and pipe lines, the centre one served as a cage-way for hoisting men and material during the operation of raising, and the third one contained a ladder way. The rest of the shaft was filled with debris during the period of shaft raising.

Jan 1, 1944

By C. R. Lattanzi, M. J. Davie, L. D. Smith, J. R. Goode

"The feasibility study determines the technical and economic viability of a project at the level of accuracy required by the owner to make an investment decision. The economic analysis usually includes calculation of the discounted cash flow. Key components of the cash flow include the project capital cost, the operating costs, construction schedule, production rate, product prices and financial details.This paper examines the feasibility study for a typical mining project. The paper discusses acceptable methods for the determination of the geological and mineable are reserves. The prerequisites for the development of suitable designs, and the calculation of capital and operating costs for the mine, processing plant, and infrastructure are reviewed. Important features of the cash flow analysis and its interpretation are presented.IntroductionThe commercial development of a mineral deposit invariably requires the investment of capital in preparing the deposit for production. It has always been necessary, therefore, to demonstrate to the investor of capital that production of a saleable commodity is both technically feasible and economically viable. The instrument typically used to make these determinations is the feasibility study .Traditionally, mining companies have been rather conservative in their approach to financial management. Until about thirty years ago, the capital required to develop a new mine was most frequently provided by internal corporate cash now and virtually all of the necessary technical and economic studies were performed 'in house'. Under these circumstances, there was little standardization as to what constituted an acceptable feasibility study, and even less as to what constituted an acceptable measure of economic viability."

Jan 1, 1991

By P. Chryssanthakis

"The aim of this study is to give a better understanding for the behaviour of the composite material RRS and the synergy effect of its components while an excavation face advances during tunneling. The lower part of the Q-system scale (i.e. Q-values of 0.4 or less for a tunnel span of 10 m) is commonly used for tunnel applications in softer rock formations. This is implied in the updated Q-chart, where the dimensioning of rock support by using Reinforced Ribs of Shotcrete (RRS) comprises the use of fiber reinforced shotcrete, S(fr), rock bolts and shotcreted beams reinforced with steel rebars. The newest part of the lower part of the Q-chart gives more information on required support compared to older Q-system versions. The RRS can be described as a three part composite material with the 3rd component being the reinforced concrete beam (with no fibers) and with steel rebars that are placed in groups and sprayed with pneumatically applied shotcrete in layers. The RRS usually have a total thickness varying between 0.20 and 0.50 m and a width between 0.25 and 0.60 m respectively. They are placed at different distances along the tunnel axis depending on the rock mass quality.In this study, an effort has been made to simulate RRS in a numerical code and to study its behaviour in a real case study. A double tube, triple lane tunnel in weathered tuff has been chosen as reference material for testing RRS under changing load. This was done by writing a FISH subroutine which is implemented in the discontinuous numerical code Universal Distinct Element Code (UDEC) that takes into account the characteristics of the reinforced beam (height, width, number of steel rebars, and spacing) in the calculation process. A rather extensive parametric study has been run by using the same key input rock mass parameters and in situ stresses (?H/?v = 0.5), and changing only the RRS key input parameters in order to simulate the changing loads during the face excavation process in tunneling. The critical part of this numerical modelling work that is under investigation in this study, is the period just after the excavation: i.e 4 hours; 16 hours after excavation; and finally after the complete curing (28 days) of the cement based components. The RRS are installed in three consecutive steps as follows: a) the application of S(fr) on the rock surface in order to even it up for the RRS application; b) radial bolting; and c) installation of the reinforcing steel rebars, that form the reinforced beam.The results show that: a) the S(fr) is almost overloaded in the first 4 hours after application in both tunnels; b) on the S(fr) and steel rebars, the stresses, axial and shear loads, as well as moments are decreasing after 16 hours and 28 days; and c) the bolts loads are increasing with time. The observed reduction in stresses and loads in both the S(fr) and reinforced beam is surprising. A possible explanation is that tunnel arches are self-supported to a lesser extent as time progresses due to redistribution of stresses in the nearby rock mass through the radial bolts. The loads in radial bolts are increasing by almost 25%. This can be seen from the overall results as we observe an increase in maximum principal stresses, displacements, shear displacements along joints and hydraulic apertures compared to tunnel baseline that has no support of any kind."

Jan 1, 2015