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By Gabka T., R. E. Gregorczyk

Work to improve employees' behavior in Głogów Copper Smelter began in 2009. It has been noted that over 90% of accidents at work are attributable to human reasons. At that time, a series of external workshops was contracted based on exercises conducted by a psychologist trainer with supervisors. Having trained about 300 employees, an observation of employee behavior at workstations was started as part of the control of risk reduction measures. As a result of the above actions, the number of accidents was reduced by half between 2009 and 2017. Currently, internal workshops for communication improvement for behavioral auditors are conducted by the employees of the Smelter. Named: "I observe, react and communicate" and aimed at further mastering the elements of the safety culture in Głogów Copper Smelter. INTRODUCTION KGHM Polska Miedz is an integrated mining and smelting copper producer based in Poland and operates on the international metal market. KGHM, with annual production of over 500 thousand tonnes of copper, holds a middle position in the world ranking of copper producers. At the same time, in Europe it is the largest producer of copper from its own ore deposits, possessing its own integrated production structure, from underground mining operations to final, top-quality product. It is also worth mentioning that KGHM Polska Miedź is one of the largest Polish exporters and, apart from copper, is a significant primary silver producer in the world. The mission of our company, consistent with its strategy, is to produce copper and silver in a safe mode for workers and the environment, with reduced environmental impact. Therefore, "Głogów" Copper Smelter operations are continuously focused on the creation of safe working conditions, and minimizing the negative effects on the environment. The company meets the requirements of an Integrated Management System comprised of the following: Quality Management System ISO 9001:2008 Environment Management System ISO 14001:2004 OHS Management System according to the requirements of BS OHSAS 18001:2007 Being a proud socially responsible company, we are aware that part of an efficient OHS system is to recognize, monitor, prevent, educate and motivate, thus continuously improve working conditions for 2500 employees of Głogów Copper Smelter &Refinery. We clearly communicate that a low accident rate is beneficial for both, staff and the company, and is of utmost importance as a part of our health policy.

Jan 1, 2019

By D. -A Tremblay

Dynamic wetting spread factors and interfacial heat transfer coefficients were determined for aluminum droplets solidifying on copper substrates, The spread factors were determined from the droplet geometries while the heat transfer coefficients were obtained by solving inverse heat conduction problems, 1-D and 2-D heat transfer models are presented and the results they provided are compared, For the latter, a sensitivity analysis was also performed to determine the influence of thermophysical properties and boundary conditions. Experiments were alternatively carried out with three different gaseous atmospheres and the dynamic wetting spread factors were correlated to the interfacial heat transfer coefficients.

Jan 1, 2005

By Brad Simser

The Brunswick No. 12 Mine is a 10 000 tonne per day producer of lead-zinc-copper-silver ore that has been in operation since 1964. Mining of this multiple lens deposit is primarily by blasthole stoping techniques between 400m and 1200m below surface. The combination of a high overall extraction ratio and high horizontal stresses (twice the vertical) has elevated the induced stress regime to levels comparable to much deeper operations. Improved mine design, and a better understanding of how the rock mass responds to mining in high stress environments is essential to improve overall recoveries and maintain safety. Seismic monitoring has proven to be an essential tool at Brunswick, for both short-term safety of workers, and long-term strategic planning. Practical case studies illustrate how seismic data, particularly source parameter information, can be used to measure the rockmass response to mining and improve mining layouts in difficult regions. As the global trend continues towards mining in deeper, higher stress environments (e.g. South African Witswatersrand Basin, Sudbury Basin) and as local extraction and thus ground stresses continue to increase in existing deposits, seismic monitoring will play a critical role in mining safely and economically.

May 1, 2003

By J. Peterson, D. D. Tannant, J. Todd

"A review of blast designs and improved blasting practices at the EkatiTM diamond mine is presented along with the results of three blast-monitoring experiments. Blast monitoring was undertaken to investigate blast damage mechanisms in the mine’s well jointed rock mass. Rock mass damage caused by production, pre-shear and wall control blasts was measured. EkatiTM uses 270 mm holes for production blasting, and 165 mm holes loaded with a decoupled charge for pre-shearing. The production blasts are loaded with bulk emulsion/ANFO blends. The mine plan involves using 30 m double benches. A 30 m high pre-shear is drilled and blasted prior to the production blasting that is done with sequential 15 m benches. This paper summarizes the monitoring equipment and data gathered to date. IntroductionThe EkatiTM diamond mine is located about 300 km northeast of Yellowknife, Northwest Territories, Canada. The mine is a joint venture of BHP Diamonds Inc. (80%) and geologists Charles Fipke and Stewart Blusson (10% each). The EkatiTM mine is Canada’s first diamond mine and BHP is the operator. The mine is accessed by air and by a winter ice road.The first kimberlite pipe mined at EkatiTM is known as the Panda pipe. The Panda open pit mine will have a total depth of 315 m (Fig. 1), a width of about 600 m, and a 50° overall slope. Initial ore production is 9000 tonnes per day.This paper summarizes the pre-strip blasting and the evolution of the production and wall control blast designs from 1997 to 2000. The initial blast designs and the rationale behind subsequent changes that took place are presented along with the techniques currently in use.The paper also briefly presents an experiment done to measure the rock mass response to blasting. Three blasts were instrumented with geophones, gas pressure sensors and time domain reflectometry cables. A brief summary of the instrumentation and data gathered is included."

Jan 1, 2001

By A. S. Moghaddam

The transportation of cemented and non-cemented mine tailings in the form of paste through pipelines is one of the main stages of paste backfill operations. For the purpose of pipeline design, the rheological properties of the materials, such as yield stress and viscosity must be determined. In this study, the yield stress of two Canadian mine tailings, for gold and copper mineralization, were studied using the vane method and slump tests. The comparison between the vane method and the slump tests indicates that the slump tests generally overestimate the yield stress values for both cemented and non-cemented paste backfills. In addition, the effects of 5% Portland cement (type 10) and mixing time on the yield stress of paste backfills were investigated. The percentage of cement as a binder agent has a direct effect on the increase of the yield stress of paste backfills. Yield stress measurements show that mixing times of 4 minutes and 12 minutes have the same effect on the yield stress values.

May 1, 2007

By Dave Bulled

"This paper describes the design of a molybdenite flotation circuit using FLEET technology applied to kinetics data produced from laboratory flotation testing and distributed across the blocks of a mine plan using a geostatistical approach. The geostatistics assigned the statistical error that arises from having a limited amount of sample testing data to the kinetics estimates of each block of ore in the mine plan. The final design then included a quantification of risk based on the accuracy of the block flotation kinetics and model fitting. The design made reference to the extensive technical reports describing earlier metallurgical test work including some pilot plant work over thirty years ago.The testing involved conducting MinnovEX Flotation Tests (MFT) to determine the rougher flotation kinetics at the specified reagent conditions and rougher-cleaner tests to determine the change in kinetics resulting from regrind. These flotation kinetics were used within FLEET simulations to create a full-scale flowsheet and plant design including recommended equipment sizes and the forecast of operating results. Measurement of detailed flotation kinetics for a wide range of samples allowed the effect of size of grind and efficiency of classification (i.e. width of size distribution) to be predicted for the whole ore body.INTRODUCTIONAdanac Moly Corp. proposed to develop a 20,000 ton per day mill to produce high quality molybdenite concentrate from its Ruby Creek resource in north-western British Columbia. It is a relatively clean low grade porphyry-monzonite occurrence containing minimal quantities of other sulphides such as chalcopyrite, pyrite, sphalerite and galena. Molybdenite occurs mainly as coarse flakes with lesser amounts of finer sized values. An optimum plant design was required to can achieve the predicted operating results, at a known level of risk, while treating the variability inherent in the ore body."

Jan 1, 2007

By S. Pinard, J. MacLellan, G. Ehlert

"The Kidd Creek concentrator is looking actively for alternative technology as means of reducing operating costs in their concentrate handling section. The use of pressure and vacuum filters to eliminate the multiple hearth dryers would have the most beneficial effect. The dryers which constitute the greater portion of concentrate drying costs consume about 35,000 m3 per day of natural gas.After the evaluation of various filters by Kidd's technology group, it was clear that, because of the low final moisture requirements of 5.5 to 6% H20, few filters were able to achieve this target economically. Lately attention has focused on the Andritz hyperbaric filter. Testwork has shown that target moistures can be achieved with an appreciable reduction in operating cost.1.0 INTRODUCTIONThe Kidd Creek Concentrator is located 28 kilometers east of Timmins, Ontario. The ore is hauled by the Kidd Creek Railroad to the concentrator where it is milled to produce separate copper and zinc concentrates.Projected concentrator feed for 1991 is 3,750,000 tonnes of ore grading 3.3 % Cu and 4.7 % Zn. The concentrate production is expected to be 458,080 dry tonnes of copper concentrate at a grade of 25.7 % Cu and 265,020 dry tonnes of zinc concentrate at a grade of 54 % Zn.Currently, there is active investigation toward cost reduction through implementation of new technologies. One area which has been extensively evaluated is the concentrate dewatering and drying section. In that section, the multiple hearth concentrate dryers represent the largest source of operating costs, mainly through maintenance and energy costs (electricity and natural gas). The replacement of these dryers with filtration technology that could produce the required concentrate dryness would represent substantial savings.In September of 1990 filtration testwork carried out by Andritz Ruthner showed that Kidd Creek copper and zinc concentrates could be filtered to below 6 and 8 % moisture respectively. In late May 1991 the laboratory testwork was duplicated on siteand one month later a 5 m2 pilot filter was tested on site. Based on the pilot test results, an economic evaluation was performed. It indicated that a reasonable payback could be realized if hyperbaric filters were to replace the existing concentrate drying section."

Jan 1, 1992

By K. Awuah-Offei

Long-term production scheduling optimization has been a challenging issue for the mining industry because of the size and complexity of the problem. The current planning algorithms have limitations in terms of addressing the stochastic variables underlying the mine planning problem. In this paper, an intelligent agent-based mine planning framework based on reinforcement learning is introduced. The long-term mine planning is modelled as a dynamic decision network. The intelligent agent interacts with the block model by means of a stochastic simulation and employs a Q-learning algorithm to learn the sequence of pushbacks that maximizes the net present value of the mining operation. The intelligent open pit simulator (IOPS) was implemented with an object-oriented design in Java®. A comparative application case study was carried out to verify and validate the models. The proposed method was used in planning an iron ore deposit and the results were compared to the Milawa scheduler used in Whittle® software. The outcome of the study demonstrates that the intelligent agent framework provides a powerful basis for addressing real-size open pit mine planning problems.

Jan 1, 2008

By J. Paraszczak

ABSTRACT Diesel-powered loaders are a pillar of modern underground metal mining. As these mines are getting deeper and hotter and the emission regulations stricter, ventilation costs are soaring, affecting production cost. Also, the cost of diesel fuel is expected to rise. These may make the use of diesel engines more problematic. Electric load-haul-dump machines (LHDs) are an interesting alternative. This paper compares features and performance of diesel and electric, cable-powered LHDs. It overviews equipment offerings, and compares, analyzes and discusses technical parameters, operational problems and cost issues. The paper gives some conclusions about electric LHD applicability and performance, indicates possible applications, and suggests further research work.

Jan 1, 2013

By Martin Stoklossa

Surface mining operations typically require run-of-mine material to be sized for further processing. In-pit crushing facilities serve this need, as stationary, semi-mobile or fully mobile units. Although in particular fully mobile crushing facilities have been in use for over half of a century, other surface mining techniques have typically been implemented. However, with increasing operating costs in truck and shovel operations, and other economic considerations, mobile crushing plants are increasingly a viable option. The operational and economical parameters in relation to the optimal surface mining equipment selection will be reviewed. The role of the mobile crushing technology within the continuous vs. noncontinuous surface mining operation and basis for mine planning are also discussed. Current applications and respective case studies will be reviewed. Key performance indicators will be presented. These performance parameters can be viewed as a tool to aid in the equipment selection process. The future direction for further research and development in mobile crushing facilities will also be presented.

May 1, 2008

By Gilles Landry, Mohamed Ourriban, Pierre Pelletier, Ahmed Bouajila, Claude Gagnon

"This paper summarizes the work carried out to develop and implement the gold cyanidation by size class (CSC) process, a novel concept aiming to reduce cyanide consumption.First, the authors will present the methodology and results of laboratory tests conducted to evaluate the response and the expected gains for the selected ore when submitted to the CSC processing. In the second part, focus will be on the methodology behind the CSC scheme development prior to its implementation at Sleeping Giant Mine concentrator. Finally, the evaluation of the modified circuit performance (operating since July 2007) will be addressed as a reduction of cyanide consumption by more than 30% and of free NaCN in the tailings by more than 50% is confirmed.INTRODUCTIONCyanidation is the most widely used technique to extract gold from hard rock ore. Its use of oxygen as an oxidant still makes it the cheapest process available after more than a hundred years. Unfortunately, gold cyanidation is known to be a slow process that requires high consumption of cyanide. In fact, cyanidation plants are designed with very long residence times (Habashi, F., 1967). Drawbacks of long cyanidation times are, among others, loss of gold in tailings, high reagent consumption and high capital investment and maintenance costs.Responding to their members expectations for the improvement of cyanidation process, COREM undertook a challenging R&D program covering the development of alternative concepts (processes) and the understanding and solving of mass transfer, gold passivation and process water chemistry problems. One of the interesting findings of the R&D program is a patent pending process technology based on the cyanidation by size class (CSC) concept (Côté and Ourriban, 2004; Gagnon and Ourriban, 2007; Gagnon and Ourriban, 2004).The new technology was tested at the Sleeping Giant concentrator, a plant member of COREM. The existing flowsheet was modified in order to implement the CSC process in the full-scale plant. The plant has been operated using the modified CSC based process since August 2007. Laboratory testwork was conducted prior to the industrial application in order to properly design the new process scheme. The modifications included:1) Reclassification of the grinding circuit product through an hydrocyclone2) Leaching of the coarse size fraction3) Addition of the fine fraction to the coarse leaching circuit tails."

Jan 1, 2009

By S. Foged

Umicore and MWH jointly demonstrated that mechanical dewatering of goethite and neutralised tailings significantly increased the life of a 30 ha tailings pond and improved the operation of the pond. Umicore operates one of its zinc production plants in Balen, Belgium. As part of its production process, an excess of neutralised and goethite¬containing tailings is generated. These have historically been stored inside large lined ponds. As the storage area of the current goethite pond was nearing its capacity, Umicore studied the potential for an extension to the existing embankment. However, this extension would face intense regulatory scrutiny, because the environmental and safety impacts were considered important. Together with Umicore, MWH suggested and investigated an alternative treatment approach consisting of mechanical dewatering prior to storage. A feasibility study was prepared where wet storage was compared to dry storage. Differences in safety risks, environmental impacts and economic costs were outlined. The study demonstrated that the life of the tailings pond could be increased from 6 to 22 years, seriously reducing the environmental impacts and safety risks at no significantly higher cost than would have been incurred with the continued operation of the existing pond. The results of the study are presented in this paper.

Jan 1, 2006

By Lionel A. York

THE first application of the cementation process took place in the year 1864, according to H. Standish Ball, when a break in the brickwork lining at a depth of 270', in one of the Rhine Preussen pits near Homburg, was sealed by running in milk of cement from the surface by means of a small hand pump. At about the same time Monsieur Portier adopted the method in France for repairing breaks in shaft lining. Figure 1 is descriptive of the Portier system and is of interest when compared with modern practice. During the next thirty years numerous attempts were made to improve the process; the general method used in sinking being to put clown bore holes outside the perimeter of the shaft for the purposes of injection. The cementation process was first recognized as an important as-set to mining in 1896 when Monsieur A. Francois developed his method of boring and injecting inside the shaft below the permanent water level. The Francois cementation process rapidly attained fame on the Continent and many shafts were sunk by its use in France, Belgium and

Jan 1, 1961

By Erwin L. Zodrow

THE AUTHORS welcome this timely opportunity to comment o n a paper by MacLean and Merner (" Room-and- Pillar Mining at the Prince Mine of the Cape Breton Development Corporation " , ClM Bulletin , June 1978, pp. 59-62) concerning the rock falls and their relation to "a change in the strata". First, however, it is proper to express our gratitude to the Cape Breton Development Corporation, and in particular to A. R. MacLean, for making it possible for us at the College of Cape Breton to start two new collections. The first of these is a fossil plant reference collection from the Sydney Coalfield, Cape Breton (Upper Carboniferous period) , and the second is a documentation of the role o f pyrite in the formation of hydrated sulphates (samples from No. 26 Colliery, 1-B Mine, and the Prince Mine at Point Aconi). Conditions of sedimentation in the deltaic-type basin of the Upper Carboniferous time in the Point Aconi area were such that on the (coal) swamp, now represented by the Stubbart (or Hub) seam , (a) clays of varying thickness and extent were deposited , (b) biogenetic sulphur was introduced and (c) flora were buried. These are essentially the culprit s that make rock falls possible and at the same time predictable over large areas (in the present mine). The clays lithified to become shale, and biogenetic sulphur combined with iron to become pyrite (marcasite?), which exists as nodules and in fine-grained layers.

Jan 1, 1978

By P. G. Rueffel

"Natural deposits of sodium sulphate are the result of normal geological processes. They are world-wide in occurrence and are found in depressions in arid and semi-arid regions. Primarily, the sodium sulphate deposits of the prairies are post-glacial and differ from rock salt deposits, which originated from connate water (fossil sea-water). Sodium sulphate on the Canadian prairies occurs as brine or in beds of intermittent and/or permanent crystals. Most beds are contaminated with various amounts of silt, sand and decayed organic matter. Soluble impurities such as chlorides, carbonates and bicarbonates of sodium and sulphates of magnesium and calcium are leached out from the soil of the surrounding run-off area of the deposits.In this paper, the origin of sodium sulphate, and the physical and chemical aspects related to the development of the Ingebrigt Lake deposit are dealt with. The problems of garnering raw material, and of refining and de-hydration are discussed, with emphasis on the Ingebrigt Lake plant. The utilization of the ""Total Power Concept"" in a sodium sulphate plant is described. The principal uses of sodium sulphate are also discussed."

Jan 1, 1968

By C. S. Maggs, L. Shwydiuk, H. W. Newcomen

The southwest wall of the Lornex Pit at the Highland Valley Copper mine has been experiencing slope displacements since 1978. Measures to manage the displacements on previous pushbacks have included stepping out, scheduling pushbacks to unload the crest of unstable areas, slope dewatering and mining with intensive slope monitoring. The final pushback on this wall is currently being mined and substantial modifications to the slope geometry could prove to be uneconomical. As a result, a campaign of surface water collection and diversion, horizontal drainhole drilling and deep well drilling is currently underway to maintain an acceptable degree of stability to the end of mine life. The various factors contributing to instability, including engineering geologic controls, rock mass strength, and surface and groundwater hydrology are presented. Movement trends, displacement magnitudes and stability analyses results are also discussed.

Jan 1, 2003

By Y. Shu, A. R. Laplante

"A laboratory Knelson Concentrator (LKC), 7.5 cm in diameter, was used to study the performance of a similar unit at plant scale, 76 cm in diameter, This paper reports on sampling and assaying procedures and problems, and discusses the performance of the plant unit.The sampling procedure was based on the extraction of large samples, 5-15 kg, and their complete processing with a LKC, to characterize size-by-size performance. The procedure yielded reliable estimates of overall gold assays, after blatant assaying errors were corrected. The plant centrifugal unit concentrate was sampled in situ, which yielded very erratic grades, with a large overall negative bias. Due to the very heterogeneous nature of the concentrate, which can be linked to flow patterns in the riffles of the unit, it is recommended that for future testwork, the concentrate be recovered separately, blended and then sampled.The plant centrifugal separator, when operated at a feed rate of 25 t/h, yielded total and free gold recoveries of 41 % and 62%, respectively. This is much lower than previously reported data, and suggest that at 25 t/h, the unit is overloaded. However, recovery did not decrease with time over a loading cycle of 2 hours, which suggests that present loading cycles, generally 1 hour, can be significantly increased. INTRODUCTIONThe study of gold gravity circuits is plagued with the statistical problem of determining the behaviour of particles occurring with a very low frequency. This problem becomes particularly acute with increasing particle size. For example, considering the average weight of a gold flake in the 840-1200 µm (20/28 mesh) to be about 5 mg (Banisi et al, 1991), a reasonably high grade feed of 10 git would require, for a relative error (standard deviation) of 10% on grade, 100 particles, that is 0.5 g of gold, or 50 kg of 840-1200 µm material (a Poisson distribution is assumed). Gy's formula, for the same data, yields a weight of 33 kg, with a shape factor of 0.1 and a size distribution factor of 0.5. Both results are comparable; to obtain the suggested weight of material, a stream containing 5% weight in the 840-1200 µm would require 1 tonne samples (of unsized material). The problem rapidly disappears with decreasing particle size. Still with Banisi et al's data, a gold flake in the 300-420 µm range averages 0.5 mg in weight; the same 10% error on grade corresponds to a weight of 5 kg. Gy's formula for the same size distribution and particle shape factors as above predicts a required mass of 1.4 kg. This would correspond to an unsized sample weight of 10 kg with 14% weight in the 300-420 µm, for example."

Jan 1, 1992

By T. Backers

In fracture mechanics modeling the rock mass is seen as the combination of rock material and fractures, which are evident on different scales. The activation and propagation of fractures dominates the stability and performance issues as well as fluid flow potential. Once fractures have started to grow into each other and create through going separations, stability of rock engineering constructions may be violated. Further, as fracture mechanics inherently incorporated a scaling law in the form of the participating fracture length, scaling issues can be more accurately mirrored. This contribution highlights some examples of application of a fracture mechanics based code on some rock mechanics issues such as numerical laboratory experiments, borehole stability or the creation of fluid pathways due to underground constructions. The results show explicitly the creation of fractures and connection of existing ones, and hence the safety and performance issues. It has to be concluded that the basis for frequent application of fracture mechanics on specialized challenges in Geomechanics is laid, and the needs for future developments are defined.

May 1, 2009

By S. Stefanizzi

In the present paper, numerical simulations based on a finite/discrete element approach are used to study deformation and failure process during standard rock mechanics. FEM/DEM modeling has been used as available with the ELFEN code, which explicitly model the transition from continuum to discontinuum. The transition is simulated through a crack nucleation and propagation that obeys to the Griffith?s failure criterion. If the fracture criterion within the intact rock (represented by FEM) is met, a new crack (represented by DEM) is initiated. Remeshing allows the fracture process through the FEM mesh to be tracked and visualized, thus contact properties can be assigned to pre-existing fractures and newly generated fractures. In this study, compressive tests of rock samples are simulated. The applicability of ELFEN to properly and efficiently simulate typical laboratory rock mechanics tests, with reference to the rock fracture process, is described in detail. In addition, an indirect validation of the extensional strain criterion proposed by Stacey (1981) is given.

May 1, 2009

By J. Lee, M. Peart, T. Bhambhani, D. R. Nagaraj, P. Riccio

"Over the past decade or so, our efforts have focused on developing the rationale and concepts for new technology in reagent development and optimization in the processing of base and precious metals. A major outcome of these efforts have been the development of a novel family of collectors, principally driven by the MAXGOLD® and Aero XD-5002® chemistries, for precious metals recovery from primary Au ores and Cu-Au ores. These chemistries was developed to be robust to address the complex mineralogy and gold deportment in these ores. Plant trial data for a variety of ores have confirmed that the both Aero® XD-5002 and MAXGOLD® chemistry are indeed versatile in recovering gold values occurring as gold alloys and/or associated with many types of sulfides, and in altering the deportment of certain sulfides (such as those bearing As, Bi etc). An overview will be given in this paper using plant data to highlight the application of these reagents.INTRODUCTIONThe subject matter of this paper is the flotation of Au containing minerals from their ores. The motivation from this manuscript comes from the several challenges faced with the processing of Au ores, as was articulated in detail by Nagaraj (2008). The biggest challenge is the complex mineralogy; most easy-to-treat free-milling ores bodies are largely depleted. The Au values are distributed in multiple host minerals such as pyrite, pyrrhotite, arsenopyrite and Cu minerals such as chalcopyrite. Au values also happen to be finely disseminated within the various minerals and are sometimes invisible in optical microscopy. Additionally, on any given day, some mines treat multiple ore types, and the mineralogy is different for each ore type.Xanthates continue to be the collectors of choice at most Au and PGM operations around the world. Most primary Au operations tend to consume very large amounts of collectors; the dosage ranges anywhere between 40-400 g/t and this has been attributed to both mineralogical and non-mineralogical factors (non-mineralogical factors include the low cost of xanthates and the mindset that “more is always better”, which is not necessarily the case). In addition, several studies now show that xanthates alone are never sufficient where multiple value minerals are present, and that supplementary collectors are required. The large doses present a substantial cost to the minerals industry, and many mining companies have expressed the need for more efficient reagents that can improve metallurgical performance at low dosages."

Jan 1, 2013