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By Oberlando Gonzalez, Claudia Serna, Andrea Alzate, Esperanza López, Maria Holuszko

This paper presents a systematic approach on gold recovery from electronic waste (e-waste) using ammonium persulfate. This process was developed as a response to the lack of hydrometallurgical systems capable of separating gold from its metallic substrate without material grinding, with minimum formation of pollutants, and achieving a lower reaction time. Computer memory boards, electronic processors and electronic pins and contacts were analyzed to determinate base metals (Ni, Fe, Cu) and gold (Au) using microwave plasma atomic emission spectroscopy (MP-AES). An aqueous commercial grade ammonium persulfate with oxygen and pressure was used to produce the persulfate anion (S2O82-) and the oxidative sulfate ion (SO42-), which partially oxidized and leached the base metals breaking the Au-Ni-Fe-Cu bonds and allowing gold to be recovered in its original non-leaching state. The influence of the oxidative parameters was evaluated using full factorial (FF) and central composite designs (CCD) with response surface methodology (RSM) and first and second order models were developed. Using RSM allowed to obtain a faster recovery of gold, minimizing the agent consumption. The findings presented suggest that optimized quantities of ammonium persulfate, oxygen, pressure, temperature and liquid solid ratio could be used for selective oxidization of the base metals and to extract more than 96% of gold from e-waste.

Jan 1, 2017

By M. Sakata, H. Hata, K. Haiki

"Most of the secondary ZnO contains fluorine (F) in large quantities. Serious troubles at the zinc electrowinning are occurred by the cathode corrosion caused by F, therefore we developed iron compounds as F-adsorbents used in zinc sulfate solution. The iron compounds can be reused repeatedly as F-adsorbents because they can adsorb/desorb F alternatively depending on the pH. Moreover, if the adsorptive capacity decreases, we can dissolve and re-synthesize them easily. In this paper, results of the experimental process which consists of five steps (leaching, F removal, Cl removal, purification and electrowinning) to produce high purity electrolytic zinc from the secondary ZnO are reported. INTRODUCTIONThe recycle material for zinc, secondary ZnO, contains large amounts of fluorine (F) and chlorine (Cl) which are halogen elements, compared with calcine obtained by roasting of zinc sulfide concentrate. The halogen elements cause adverse effects in the zinc electrowinning process and particularly F causes corrosion of aluminum cathode plates and eventually sticking of electrodeposited zinc and therefore, F concentration in the electrolyte is controlled to less than approximately 20 mg/L. There are several methods to remove F from secondary ZnO, such as leaching with Na2CO3 or NaOH solution [1-2], vaporizing in a rotary kiln by roasting at 1000 to 1200 degrees C [3-4]. However, up to the moment, there has been no effective method for the removal of F from zinc electrolyte and it has been a big obstacle in treatment of secondary ZnO in electrowining of zinc.The authors, therefore, developed an iron compound used as an adsorbent of F in the zinc sulfate solution [5-8]. This iron compound is Akaganeite that has a crystalline structure of ß-FeOOH. Akaganeite is synthesized by neutralization of FeCl3 solution containing a small amount of SO4 2- with Na2CO3. When this iron compound is immersed into weak acidified zinc sulfate solution, F in the solution is adsorbed. F adsorbed in the iron compound is removed by immersing in weak alkaline solution and the iron compound can be reused as F adsorbent."

Jan 1, 2012

By Ross D. Lawrence

"IntroductionOne of the methods used when applying the Market Approach to the valuation of a mineral exploration property is known as the Joint Venture Terms Method. The method is particularly useful for the valuation of early-stage exploration properties. Unfortunately, valuators are sometimes imprecise in distinguishing between joint ventures and other types of transactions.At VALMIN '94, Bob Appleyard1 proposed that, if one ignores the time value of money and the probability that the expenditure commitment would be completed, the formula to calculate the value ($V) of a property owned by Party B at the time of the deal would be:$V =$Ex (100-1%)/1% (Formula I)Where $V is the value of the property2Where $E is the expenditure commitment by Party A, andWhere 1% is the percentage property interest to be earned by party A.At PACRIM '99, Mike Lawrence and Graham Dewar3 referred to this method and suggested that there was an ambiguity in the formula proposed by Appleyard in 1994. Mike Lawrence referred to this problem again at VALMIN '014•Lawrence & Dewar suggested that the formulashould be:$V100% =($Ell%) x 100 (Formula 2)Where $V 100% is the value of the entire property,Where $E is the expenditure commitment by Party B, andWhere 1% is the percentage property interest to be earned by Party B.Subsequently, there has been considerable discussion in the AuslMM Bulletin5. In my view, this discussion has touched on a number of useful points, but generally has overlooked one essential element needed for an understanding of the correct application of the N Terms method. That essential element is to understand that there is a material difference between (1) entering into a joint venture arrangement and (2) the purchase of an interest in a property."

Jan 1, 2003

By J. R. Parga T., P. J. Valenzuela G., L. S. Quiroz C., P. Guerrero G., J. L. Valenzuela G.

"Pressure oxidation is a hydrometallurgical pre-treatment process used for refractory auriferous sulfide ores. In the work reported here, a refractory pyrite ore sample from the Mulatos mine, Sonora, Mexico was characterized. The sample was subjected to atmospheric cyanidation, and simultaneous pressure oxidation and cyanidation under different conditions of temperature and pressure. The mineral characterization showed that pyrite is the main species in the sample. The gold extraction tests were carried out in an 1-litre Parr pressure vessel at a temperature range of 90-120 °C, a pressure range of 87.5–150 psi (using oxygen), and a cyanide concentration range of 0.8-1.2 %wt. The gold extraction results showed that diffusion through product layer is the rate controlling step.INTRODUCTIONIn hydrometallurgical precious metals extraction, the most commonly used process is sodium cyanide leaching, as shown in the next reactions involving gold and silver (Fleming, 1992; Marsden & House, 2006):4Au(s) + 8CN- (aq) + 2H2O(l) + O2(g) ? 4Au(CN)2 -(aq) + 4OH- (aq) (1)4Ag(s) + 8CN- (aq) + 2H2O(l) + O2(g) ? 4Ag(CN)2 (aq) + 4OH- (aq) (2)“Refractory” auriferous ores or concentrates often contain gold and silver locked inside other sulfide phases such as pyrite. This renders cyanidation ineffective, so oxidative pre-treatments are used in order to convert the sulfide minerals into porous oxide minerals and make cyanidation work more efficiently. Reaction (3) shows the oxidation of pyrite in alkaline media and the formation of the porous oxide (Koslides & Ciminelli, 1992):2FeS2 (s) + 15/2 O2 (g) + 8H- (aq) ? Fe2O3 (s) + 4SO4 2- (aq) +4H2O(l) (3)A process of simultaneous chemical pressure oxidation and cyanide leaching for a flotation concentrate produced in the mine of Bacís, Durango, in Mexico, gave good results (Parga et al., 2007). In the study presented here, it was intended to use this treatment for a refractory pyrite mineral from the Mulatos mine in Sahuaripa, Sonora, Mexico, modifying variables such as pressure, temperature and cyanide concentration to find the best conditions for oxidation and leaching."

Jan 1, 2012

By Trang T. Tran-Valade

With continually increasing energy prices, the mining industry has been on the search for new energy optimization technologies. One area that is garnering great interest is Ventilation-On-Demand (VOD). Prior to selection and implementation of a VOD solution, it is important to fully understand the mine requirements as well as its existing infrastructure and the interaction with other automation in place at the site. With this information, a business case can be generated including a cost versus benefit comparison for various control strategies. To assist in the development of the business case, this paper will identify key areas for consideration and recommendations for an implementation plan.

May 1, 2013

By F. Bélanger, P. Fournier

"The Detour Lake Mine is located in northeastern, Ontario, approximately 300 kilometres northeast of Timmins and 185 kilometres by road northeast of Cochrane. The mill processes a free milling gold ore using both gravity circuit and leaching with a state-of-the-art carbon-in-pulp circuit. The grinding circuit is a conventional SABC circuit with a tonnage of 60,000 mtpd at 95 microns with the addition of pre-crushing upstream of the SAG mill. This paper will highlight the impact of a pre-detox thickener on the cyanide consumption in a gold leach/CIP circuit. While the pre-detox thickener was built with the intention to reduce the cyanide consumption, the opposite effect was observed when the thickener was put temporarily offline. It will show how the re-circulation of a dilute stream of cyanide into the grinding circuit, increases the overall slurry temperature and by the same effect, increases the dissolved copper concentration in the leaching circuit which subsequently increases the cyanide consumption.INTRODUCTION The processing plant is using a conventional gravity, cyanidation and carbon-in-pulp processing facility currently operating at approximately 60,000 mtpd. The comminution process is based on two parallel SABC lines, each fed from an individual reclaim tunnel. Cyclone overflow from the two ball mill circuits is directed to the pre-leach thickener. The thickener underflow is distributed to two (4) parallel leach trains of 5 tanks each, each leach train providing 29 hours of leaching retention time. After leaching, the gold in solution is recovered in a carrousel CIP circuit consisting of two (2) parallel leach trains. Tails from each CIP circuit are combined and sent to a pre-detox thickener to recover as much cyanide solution as is economically practical before being diluted and fed to a cyanide destruction system. The detoxified tailings are sent to the Tailing Management Area (TMA) for settling and retention. The excess water from the TMA is reclaimed and recycled back to the processing plant. The pre-detox thickener is designed to thicken the leach tails slurry to 64% solids with the overflow being recycled to the grinding circuit. It is estimated that the pre-detox thickener will recover 60% of the cyanide in the CIP discharge solution. It is also designed to reduce the SO2 consumption by diluting the thickener U/F to 50% solid using the tailing pond reclaim water which has very little cyanide content."

Jan 1, 2019

By George C. Riley

While the general public has little knowledge or appreciation of either the mining or explosives industries, modern civilization could not exist without them. They are to-day the key industries of the world. In common with many other industries the demands of the war greatly stimulated the production of explosives and the research necessary for such increased production. Production, supplies of all kinds, munitions of all sorts and then again production, practically regardless of cost, were the first reactions to the imperative war demands of the governments of the world. Following the war there was a momentary lull, then first gradually but with quickening acceleration, particularly in a recent slump, the situation was reversed - cost and not production was the dominant factor. Some industries readjusted themselves very rapidly and completely. Others were but partially re-adjusted and others again only carried out the process of deflation in certain departments. The explosives industry in Canada has not only thoroughly deflated itself, but has learned to apply to commercial disruptive explosives, the best, most efficient and most economical methods of manufacturing developed during the war. As a consequence, miners have at their disposal to-day better and more different kinds of compressed energy than ever before in all the long history of mining.

Jan 1, 1925

"Previous studies on the effect of polytetrafluoroethylene addition during chalcopyrite leaching under conditions similar to those of the CESL process have reportedly resulted in complete leaching of the copper without the need for grinding of the flotation concentrate. However, these studies employed a pulp density of only 1% and finely sized PTFE particle that required the use of high impeller rotation rates that are not sustainable at an industrial scale. This paper explores the performance of 1/8” diameter balls of PTFE, that readily sink in the leach solution, on leaching performance at impeller rotation rates and ore pulp densities more typical of industrial practice. Possible routes for separation of the PTFE from the leach residues and their regeneration for recycling back to the leaching stage are also discussed.INTRODUCTIONChalcopyrite, CuFeS2, is the most important primary sulfide source of copper. While pyrometallurgical smelting of chalcopyrite remains the only widely accepted route for treating chalcopyrite concentrates, several medium temperature pressure oxidation processes are claimed to be economically feasible (Muir & McDonald, 2007). Among these, the CESL process has been extensively piloted and demonstrated at the prototype scale. The overall process chemistry of the CESL process may be summarized as follows (Marsden & Wilmot, 2007):4CuFeS2 + 5O2 + 4H2SO4 ? 4Cu2+ + 4SO4 2- + 2Fe2O3 + 8S0 + 4H2O (1)4CuFeS2 + 17O2 + 2H2SO4 ? 4Cu2+ + 10SO4 2- + 4Fe3+ + 2H2O (2)Though copper extractions of over 95 % are achieved, the CESL process requires extra grinding of the concentrate (Muir & McDonald, 2007). This is because leaching is conducted at temperatures around 150 °C, where the elemental sulfur produced is molten and grows thicker on the mineral’s surfaces as oxidation progresses, creating an increasingly large barrier for mass transport. In addition, the sulfur is sticky and tends to cause the mineral particles to agglomerate, effectively halting further oxidation of the particles that end up embedded within the agglomerates. Consumable surfactants, such as ligin sulfonate, have been shown to be effective at avoiding agglomeration, but they are not effective at removing sulfur from the minerals. Hence, their use does not obviate the need for fine grinding."

Jan 1, 2012

By M. Grenon

OBJECTIVE ?To link slope stability and financial assessment algorithms into a commercially available mine design software packages to provide integrated global pit optimization.

May 1, 2013

By D. Gagnon, B. Dubreuil, J. -M. Lalancette, D. Lemieux

"The pilot scale extraction of gold with a brine of sodium chloride and bromide at 30 ºC, in an atmosphere of halogens at atmospheric pressure, leads to gold chloride with 95% gold recovery. The use of dilute hypohalites (1 – 3%) instead of free halogens, which is strong enough to oxidize hypobromite into bromine, simplifies the recycling of halogens by electrolysis. With the recycling of water and brine, the regeneration of hypohalites and with a barren solid that is inert and stable. The Nichromet Extraction technology is a green process where the protection of the environment does not preclude profitability.INTRODUCTIONThe use of chlorides for the extraction of gold predates the discovery of chlorine by Scheele in 1774. In the Plattner chlorination process (Christy, 1888) sodium chloride and copper salt were used to promote gold recovery by amalgamation with variable results. The introduction of the cyanide technology in 1889 has led to a general use of this method for gold recovery. In the last decades, a return to chlorine in extractive processes has been noted (Ferron, Fleming, O’Kane and Dreisinger, 2002), mostly because of environmental pressure for non-cyanide technologies.This paper describes a closed circuit proprietary approach (Lalancette, 2009) developed and piloted by Nichromet Extraction Inc. for the extraction of precious metals with recycled mixed halogens; chlorine and bromine from hypohalites. DESCRIPTION OF THE PROCESSThe attack of gold by halogens can vary significantly. The rate of reaction is substantially higher with bromine (Br2) than with chlorine (Cl2), by a factor of more than five, although the oxidation potential of Cl2 is higher than Br2, as shown in Equation 1a and Equation 1b (Latimer, 1952)."

Jan 1, 2012

By C. Barlow

The White Mesa Mill (WMM) began operating in 1980 as a regional processing mill owned by Energy Fuels Nuclear, Inc. Original mill design was for 656,000 Mtpa of U ore producing 1,250 tpa of U3O8 concentrate. Sustained market downturns delayed facility upgrades during most of the 1980s and 1990s. The new owners of WMM, Denison Mines, elected to advance the regional mill concept in 2007 with the modernization of these facilities. This paper details decisions taken and lessons learned in plant redesign, choice of leading-edge instrumentation, motor controls and process automation, training and some example benefits from process control stabilization.

Aug 1, 2010

By B. D. Frater

Geometallurgy or process mineralogy is an integrated part of the scoping, pre-feasibility and feasibility phases of a project as well as an existing operation. A well planned and executed geometallurgical testwork programme during the early phases, will result in a more accurate decision making process during the selection phase. The better this integration between geological, mining and metallurgical data gathering during the early phases, the more accurate the predictions will be with regards to throughput and recovery during modelling and process and equipment selection in later phases. This will enhance and improve the mine and process scheduling and subsequent NPV. Geometallurgical information gathered during plant operation and testwork can for example be used to statistically model the pit optimisation, plant throughput and optimise the plant for certain feed material (characterised by statistical models).

Jan 1, 2017

By Micheal Mayhew

"This presentation will provide information and facts on the drilling and blasting within a frozen shaft using a full face methodology. The Author will briefly discuss the history of the sinking of a frozen shaft in Saskatchewan, the challenges faced, and the solutions delivered. The case study will include a discussion describing the series of events that led to the decision to adopt a full face blasting methodology. The case study will describe key elements such as the drilling pattern, hole size and depth, explosive products and blast vibration monitoring systems. INTRODUCTION During the shaft sinking process at a project in Saskatchewan, it was recommended that a drill & blast benching method be utilized. Following an evaluation of the site, the geological conditions proved amenable to a ground freezing system that would stabilize the rock formations and prevent inflows of water into the shaft (Table 1).The shaft sinking process required that the ground surrounding the proposed shaft be frozen by drilling several near vertical freeze holes radially at a specified distance from the edge of the proposed shaft excavation. This would allow for the circulation of a chilled brine solution down the freeze holes. The circulation is achieved by means of stainless steel pipes encased within each other to provide a delivery and return conduit for the brine (Figure 1). As a result of this periphery freezing, the ground within the shaft excavation is also predominately frozen.The finished internal diameter of the shaft was 5.5 meters with varying overall excavation diameters to facilitate different liner designs used for various rock formations. For the most part, the fully unsupported/unfinished excavated diameter of the shaft varied from 6.9 meters to 8.5 meters depending on the zone excavated. The challenge was that the proposed drill and blast excavation method posed a significant risk due to the potential to damage the freeze wall and pipes during blasting. The decision was made to commence the excavation through mechanical excavation (i.e. chip and cut) methods then change to a drill and blast method utilizing a split benching design. Through this process, the team encountered five main challenges: 1. Sub-horizontal geological structures and slips that affected blasting performance and sometimes resulted in a sloping surface to the top of the bench; 2. Confined working area for the crews and equipment due to a limited work area as a result of utilizing the split (half) bench drill and blast methodology; 3. Mechanical chipping and cutting of the one-meter buffer zone to the final shaft diameter; 4. Crews working on extremely slippery uneven frozen rock and ice surfaces, which affected safety and efficiency; 5. Establishing a blast vibration monitoring system to determine if the drilling and blasting of the full shaft excavation diameter were acceptable."

Jan 1, 2019

By X. -G. Chen, A. M. Nabawy

A novel manufacturing technique for Al-TiB2 nanocomposites has been developed, in which TiB2 nanoparticles were incorporated into liquid aluminum under mechanical stirring and in the presence of a K2ZrF6 flux-wetting agent. The microstructure of the Al-TiB2 nanocomposites was studied using scanning electron microscopy (SEM) and transmission electron microscopy (TEM). The microhardness test was conducted to investigate the strengthening effect of TiB2 on the aluminum matrix. The TiB2 nanoparticles were successfully incorporated and dispersed into the aluminum melt in the presence of the K2ZrF6 flux during mechanical stirring. In the as-cast microstructure, most nanoparticles segregated in the interdendritic regions as a result of particle redistribution during solidification. The addition of 2, 5 and 10 wt% TiB2 nanoparticles increased the composite hardness by 60, 90 and 150% respectively when compared to commercial purity aluminum, indicating that the TiB2 nanoparticles imposed a significant strengthening effect on the aluminum matrix. The role of K2ZrF6 flux agent on the incorporation of nanoparticles into aluminum melt was discussed.

Jan 1, 2014

By H. Henao, E. Kohnenkamp, L. Rojas

The slags generated in the copper conversion process are mainly composed of Cu2O-Fe2O3-SiO2 with the compounds CaO and Al2O3 in concentrations up to 10 wt%. The present work contributes to the knowledge of the conversion process, generating experimental data for the diagrams of phases Cu2O-Fe2O3- SiO2-Al2O3 and Cu2O-Fe2O3-SiO2-CaO. The experiments were carried out in a tubular furnace at temperatures of 1150°C and 1200°C under a condition of saturation with tridymite. Once the equilibrium was reached, the samples were fast quenched in water. The phases in the samples were observed in a Scanning Electron Microscope (SEM) and the elemental composition of the phases analyzed by means of Energy-dispersive X-ray spectroscopy (EDS) detectors. The addition of Al2O3 and CaO into the Cu2O-Fe2O3- SiO2 system resulted in an appreciable displacement of the liquidus lines, corresponding to an expansion of the liquid in the tridymite primary phase field. INTRODUCTION The conversion process of copper consists of oxidizing Fe and S present in mattes with 60 to 65wt% Cu (up to 16wt% Fe). This process is separated into two stages: 1- Slag blowing: Fe is oxidized, the addition of a flux such as SiO2 (silica) forms a second liquid phase called slag, this phase is removed from the furnace; 2-Copper blowing: in this stage, the matte with low concentration of Fe (<1% w/w) is again oxidized to remove S as SO2 in the exhaust gases. The product of these two processes is the formation of Copper Blister (> 98wt% Cu) (Schlesinger et al., 2011). Unlike the aforementioned conventional processes, in most Chilean smelters, conversion Peirce- Smith Furnaces are loaded with high copper matte (70–75wt% Cu) (Henao & Moyano, 2016). Due to the low concentration of iron in these matte (up to 5wt%), it is considered that separating the conversion process in two stages is neither operational nor environmentally convenient. Therefore, the process is carried out in a single stage (Pradena et al., 2006), generating a slag phase with a high percentage of Cu2O (wt%Cu>40).

Jan 1, 2019

By T S. Cochrane

A variety of practical tests have been tried in the mining industry to determine the support properties of mine roof bolts. Such tests include torsion wrench readings, hammer blow tests and the use of plates with built-in tension indicators. This paper gives some details of pull tests with a hydraulic jack, as carried out in coal mines of western Canada. M A~Y criteria exist for establishing the efficiency of roof bolting, and the one adopted de-pends on the nature of the problem and on the individual leanings of the investigators. Consequently, sufficient studies have not been made from a common viewpoint or under similar geological and techno-logical conditions to compare these methods of evaluation. However, regardless of whether certain tests provide sufficient factual data to form a basis for theoretical research on bolting or not, many of these tests can and do provide a positive means of judging the suitability of chosen roof bolting techniques and equipment under definite geological conditions.

Jan 1, 1963

By Michael Loken, April Montera

"The process of blending consistent and predictable feed for the roasters is critical to optimizing product quality and controlling production costs. To assist in this process, Online PGNAA Slurry Analyzers have been installed and evaluated at a Sudbury Ontario smelter. They are used to measure 100% of the full process stream to obtain elemental composition data in real time. This paper explains the application details at Sudbury and the benefits achieved from their Full Stream, Elemental Slurry Analyzers.INTRODUCTIONSudbury Integrated Nickel Operations (Sudbury INO), a Glencore Company consists of two underground nickel/copper mines in Sudbury, Nickel Rim South and, Fraser, along with the Strathcona mill and the Sudbury smelter. The facilities are spread throughout the 60 kilometer-long, oval-shaped geological formation known as the Sudbury Basin (Figure 1). The company has been mining nickel-copper ores in the Sudbury area since 1929. Nickel and copper are the primary focus of operations at Nickel Rim South and Fraser. Other base metals such as cobalt and precious metals are also produced."

Jan 1, 2016

By Gurupraveen Ramanathan

Shallow-dipping orebodies (45°–60° dip) present significant challenges for underground longhole stoping, particularly due to limited gravity flow, drill misalignment near the footwall, increased potential for dilution from hanging wall overbreak, and deviation from drill string sagging at shallow angles. These geometrical and operational constraints often lead to unblasted ore wedges and excessive overbreak when standard practices are applied. This study develops a structured, data-driven framework to optimize longhole stoping in such conditions by integrating five key focus areas: (1) mining method selection, (2) stope and drift design, (3) drill and blast optimization, (4) equipment selection, and (5) performance evaluation. The suitability of longitudinal longhole stoping was validated for narrow vein geometries using orebody analytics and prior case studies. Geometric modeling was employed to determine optimal stope height (15–20 m) and stope width (1.5–2.5 m), with drift widths adjusted based on orebody dip to maintain drill access and minimize unbroken ore. Drill and blast design emphasized sequential blasting over multi-ring methods to improve void control, reduce wall damage, and enhance fragmentation. Specific burden and spacing guidelines (e.g., 1.8–2.3 m burden; 2.5–3.0 m spacing) were applied based on stope width and dip angle, and patterns were selected based on geometry—dice-5 for narrow stopes, and 3-hole rows for wider stopes. Drill rigs—including top hammer and ITH models from Sandvik, Epiroc, Boart Longyear and CMAC-Thyssen—were evaluated for compatibility with drift dimensions and alignment accuracy. Equipment was matched based on dip, drift width, and the ability to access the footwall at shallow angles. Sensitivity analysis quantified how changes in geometry, drilling accuracy, and blast design affect ore recovery, dilution, and operating cash flow. Results show that tailored drift and stope geometries, combined with proper blasting and rig selection, can boost ore recovery to over 95% and reduce dilution to below 5% in optimal conditions. This methodology offers a practical and economically optimized solution for shallow-dip mining scenarios.

May 4, 2025

By Victor F. Hollister

Elemental sulphur occurs as aformational sedimentary deposit in Upper Devonian Wabamun evaporites near Coronation, Alberta. The deposit developed in a stratigraphic jacies-change lrap; it is assumed that petroleum reacted with anhydrite to form sulphur, calcite, hydrogen sulphide and other reaction products. The deposit contains about 362,800 tonnes of sulphur and lies at a depth of 1200 m. The Coronation deposit is expected to be geologically typical of such other native sulphur deposits as may be found in Wabamun Group sediments. Widespread sour gas sulphur production f rom the Wabamun suggests the possibility that additional native sulphur deposits will ·be found.

Jan 1, 1984

By B. Chambers, M. Yunovich

"Corrosion monitoring is commonly used in a variety of industries as a means of effective corrosion control programs to extend the life of the asset and ensure safe operation of the facilities. The majority of the monitoring tools have limitations and tradeoffs – e.g., sacrificing the longevity of the probe for higher measurement resolution. Linear polarization resistance (LPR) probes have the advantage of providing the operator with accurate, real-time corrosion rate data; however, the probes can only be deployed in conditions where the measured environment possesses certain minimal conductivity. In many instances, utilization of LPR probes in conditions with intermittent wetting (for example, conditions where the temperature fluctuates above and below the dew point) presents a challenge. To address this limitation, a special-design cooled corrosion probe has been developed, commercialized and integrated into the distributed control system in the plants. In addition to high-accuracy measurement of corrosion rates, the probe allows for monitoring of pitting corrosion. The most innovative feature of the probe is the cooling circuit, which brings the temperature of the sensors below the dew point, forcing precipitation of moisture and thus permitting corrosion measurements. Additionally, the probe was found to be well-suited for monitoring corrosion in crude distillation unit (CDU) overhead space, where condensation of hydrochloric acid may take place at relatively high temperatures, which would require relatively moderate amount of cooling to be imparted to the probe sensors. This paper presents laboratory-based data from a test simulating CDU refinery environment. A case study is presented to demonstrate how the cooling probe can be effectively used in the plants for corrosion monitoring.INTRODUCTIONConventional corrosion-monitoring techniques such as weight-loss coupons to assess corrosion damage are commonly used in plants1. However, the use of coupons is not only labor-intensive, but also does not deliver real-time corrosion data, thus effectively serving as a “backwards-looking” monitoring tool. Linear polarization resistance (LPR) probes have the advantage of providing the operator with accurate, real-time corrosion rate data; however, the probes can only be deployed in conditions where the measured environment possesses certain minimal conductivity."

Jan 1, 2012