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By J. J. Robertson, D. M. Polizzotti

Cement is used as an established agglomerating agent in gold heap leaching operations to maintain heap permeability and prevent migration of fines. In addition to agglomeration, cement provides protective alkalinity control. However, the amount of cement needed for agglomeration is usually very different from the amount necessary for pH control. As a result, cement is usually added in excess. Overdose of cement can result in reduced gold recovery and/or lengthy detoxification. Betz Laboratories has identified polymeric agents that provide the benefits of cement without negatively impacting gold recovery when used alone or with lime. Independent laboratory tests also strongly suggest that the polymer/lime system can significantly reduce heap detoxification time.

Jan 1, 1992

By Nick Katsikaros, A H. Gray

Gold concentrates are produced at many mines by a variety of centrifugal and gravity machines, such as Knelson and Falcon concentrators, however the subsequent recovery of gold is relatively inefficient and subject to security problems. The InLine Leach Reactor (ILR) offers a new method to recover over 95% of gold in concentrates using a simple, secure leaching process. The InLine Leach Reactor is of particular benefit in treating slow or difficult leaching sulphide concentrates and its performance can be reliably predicted from simple laboratory tests. The InLine Leach Reactor is now successfully operating in two gold mines providing benefits of high gold recovery, low operating costs, high security and ease of operation.

Jan 1, 1999

By K. A. Prisbrey

Mineral carbonation is the reaction of carbon dioxide with mineral silicates to form solid, carbonate compounds. The carbonate ion holds carbon in a low energy state – lower than in carbon dioxide. Thus the reaction of carbon dioxide to produce mineral carbonates captures the carbon in a stable, solid form. The University of Utah is investigating the carbonation of silicate minerals by mechanochemical processing, which has the potential of being much less expensive than other methods. Tests to date have shown that a marked change takes place in the lizardite variety of serpentine after 15 to 20 minutes of intense grinding in the presence of gaseous CO2. Further testing is in progress.

Jan 1, 2004

By J. Koinig, K. Mussger, St. Reischl

Jet grouting is a viable supplementary measure for NATM tunneling in special softground conditions. In particular it is used as a type of forepoling to maintain face stability and to provide initial support for large openings until the shotcrete lining reaches sufficient strength. For this purpose cement grout is injected under high pressure through horizontal pipes outside of the tunnel perimeter and thoroughly mixed with the soil producing 1 to 3-foot diameter cylindrical beams. Excavation of the tunnel then proceeds beneath this umbrella of 40 to 50 foot long beams. Design and construction experiences from recent projects are presented for tunnels diameters between 30 and 50 feet in fill and alluvial and colluvial deposits.

Jan 1, 1987

By Paul Chamberlin

Although performing a heap leach test is inherently simple, a lot of information beyond extraction, leach time, and reagent consumption can be obtained from a well designed test program. The following is a list of information to look for and acquire when performing column tests and field heap leach tests.

Jan 1, 1998

By J. Bilant, J. Donnelly

"A spatially-discretized heap leach simulation model was developed within the GoldSimTM modeling framework for the Coeur Rochester Mine, Nevada. The model is being evaluated by Rochester as a means of improving forecasting of silver and gold production on an actively loaded valley fill heap leach facility. Optimization of future heap leach operations and crusher product size is also being evaluated with this software. Simulation modeling incorporates metallurgical data such as mine planning, particle size distribution of placed ore, stacking plans, pad geometry, and leaching dynamics. The heap leach model is represented as a discretized matrix with lifts providing vertical separation to form 19,840 spatially discrete cells. Each cell is quantified with regard to ore properties and metal content. The model tracks solution and cyanide-complexed silver and gold within each cell as a function of time based on pad operations, unsaturated flow properties, and kinetic rates. Practical heap leach management and daily operation will be discussed as it relates to the simulation and forecasting model. INTRODUCTION The Rochester mine, and associated valley fill heap leach facilities, is an open-pit silver and gold mine located in Pershing County, Nevada. In 2015 Rochester produced approximately 4.6M troy oz silver (Ag) and 52K troy oz gold (Au) through combined cyanide heap leaching of the Stage II, III and IV heap leach pads with recovery via a Merrill Crowe process plant. Rochester is actively stacking and leaching ore on the Stage III Heap Leach Pad. The Stage III pad contains approximately 70M tons of ore in a valley fill configuration and is forecasted to receive an additional 20M tons of ore by the life-of-pad in 2019. The ultimate height of the pad is permitted to be 400 ft above the liner and solution collection system. Approximately 80 percent of the ore loaded on the Stage III pad is produced from tertiary crushing split from two crushing circuits that produce a nominal 80 percent passing 3/8” product. The remaining material consists of run-of-mine (ROM) from off loads of re-mined waste rock stockpiles. Metallurgical testing of the ore indicates low variability in geologic properties and high variability of both Ag and Au extraction rates for the various particle size distributions. Historically the valley fill configuration provided metal forecasting challenges that could not be predicted as the leach pads grew. Forecasting metal production and leach optimization of the Stage III pad became challenging due to the extreme variation in heap leach pad depth, application of available barren leach solution and the variability of Ag and Au extraction rates for the various ore particle size distributions. To assist with understanding and optimizing metal production forecasting with these factors for the Stage III pad, Hatch developed a spatially-discretized heap leach simulation model using GoldsimTM dynamic simulation software. The model incorporates metallurgical data such as mine planning (tonnages and grade), particle size distribution, ore crushing, ore stacking plan, pad geometry, and both leaching dynamics and kinetics to forecast metal production from a consolidated mine plan."

Jan 1, 2017

By Kenneth L. Clifford

We have heard a lot during the recession about the mining industry now competing in a world market for sales of such commodities as copper, lead, zinc, molybdenum, iron ore, and phosphate. If the US mining industry is to survive, do we not have to reduce direct operating costs to a minimum? How significant is labor cost compared to consumables, electricity, fuel, maintenance, materials, and transportation? Ore grade is ultimately the dominant factor; but can we offset its dominance by improving productivity? Is the US mining industry perhaps only 25-30% as productive as it might be? What prevents the industry from being more productive? Can we retrofit older plants to be more productive or does it take a new plant with the largest equipment available for each unit operation? Present Productivity A copper industry study a few years ago covering 13 US copper concentrators ranging from about 10-45 kt/d (11,000-50,000 stpd) showed that productivity did not vary with capacity above 18 kt/d (20,000 stpd). Operating productivity varied significantly, however, from about 180-635 t/d (200-700 stpd) per operating hourly worker in most units, but one operation had 900 kt/d (1,000 stpd). Concentrators with more than 545 t/d (600 stpd) per hourly worker were in the 23 kt/d (25,000 stpd) or more category. Maintenance productivity covered about the same range, but most concentrators were in the 270¬635 t/d (300-700 stpd) per hourly maintenance worker. Again, one company stood well above others by exceeding 1 kt/d (1,100 stpd). Total hourly productivity ranged from 125-280 t/d (140-310 stpd) per hourly worker for most concentrators. The company with the most productive operation, however, was above 455 t/d (500 stpd). Salaried worker productivity for most operations was in the 725 t/d to 1.2 kt/d (800-1,300 stpd) per salaried worker range. The best operation was 2.2 kt/d (2,400 stpd). The end result is, of course, tons per day per total personnel. Most operations fell in the 110-220 t/d (120-240 stpd) range with the best operation at 400 t/d (440 stpd). This is the productivity definition that is most directly related to unit costs. Many of these operations have made some productivity progress

Jan 11, 1983

By Kwadwo O. Adutwum, George M. Potter, Todd J. Harvey

Ashanti Goldfields Company (AGC) is the largest gold mining company in Western Africa. Their jewel operation is a highly inte­grated set of mines and plants situated in Obuasi, in the Ashanti region of Ghana. This operation is responsible for mining 5.5 mil­lion tonnes of ore at 4.6 git annually from a series of open pits and underground mines. The sulphide ore is highly refractory with untreated gold recoveries in the range of 30-40%. The milling facilities are quite varied and include an oxide treatment plant (OTP), a heap leach, a sulphide roaster circuit (PTP), a tailings treahnent plant (TIP) and the most recent, a sulphide treatment plant (STP). The sulphide treatment plant, the world's largest stirred tank biooxidation circuit, was commissioned in early 1994. This plant treats approximately 800 tonnes of refractory gold con­centrate per day with overall recoveries of75-85o/o(ll, Biooxidation was selected for its ease of operation and less exotic maintenance requirements as well as its ability to achieve comparable gold recoveries. Unfortunately, despite stirred tank biooxidation being one of the lower cost methods of treating refractory ores, it is still too costly for low grade refractory materials. GeoBiotics, Inc. has developed a proprietary technology that allows the economic treatment of lower grade refractory gold ores through heap biooxidation. AGC has a significant resource of low grade refractory sulphide ore. This ore is too low grade to warrant BIOX™ sulphide treatment and its oxide plant recov­ery is low at 35-45%. GeoBiotics and AGC have completed pre­liminary testing and economic analysis of heap biooxidation for this material. Additionally, the feasibility of producing a gravity concentrate from the tailings of the sulphide treatment plant has been explored. This concentrate would be coated onto the low grade sulphide and biooxidized concurrently. Gold recoveries for the biooxidized low grade sulphide have increased to 75% and the gravity recovery from the STP tailings has achieved 80% uti­lizing a Falcon concentrator. The preliminary economics of the proposed full scale plant are very favourable with a pre-tax IRR of 233% and a payback period of 6 months. This paper details the biooxidation testing, flow sheet devel­opment and preliminary economics of employing heap biooxi­dation for low grade refractory ore treatment at Ashanti Goldfields' Obuasi operation.

Jan 1, 1998

By John A. Lombardi

The solutions used in the heap biooxidative treatment of sulfidic refractory Au ores and concentrates generally load free mineral acid and metals and, in time, become bacteria activity inhibiting due to increased acid and/or metals concentrations. The metals­acid loaded, bacteria action inhibiting, solution must then be: (a) processed, prior to re-use, to reduce its metals and/or free acid mineral acid components to bacteria activity compatible levels; or (b) neutralized prior to discharge. The most common solution processing method to adjust acid-metals contents is liming to effect the precipitation of dissolved metals and neutralization of acid sufficient for solution discharge or re-use. The lime treated solution is high in suspended, colloidal precipitates and requires clarification before reuse. Most often the "bleedstream" of biooxidation solution that is lime treated is discharged and req­uisite volumes of fresh water added to the process inventory. Depending on the level of bacterial action in a heap and the acid consumption characteristics of the ore/concentrate being biooxi­dized, the bleedstream taken from the circulating, irrigation on­off, solution, can be large (e.g., 10%-20% and sometimes as much as 50% of the daily solution flow requires treatment-dis­charge). The lime charge for bleedstream treatment is commen­surately large and represents a major operating cost of the bioox­idation process. The patented HW Process Technologies (HWPT) Engineered Membrane Separation (EMS™) method is an alternative biooxi­dation bleedstream treatment method. EMS™ processing removes the multivalent base metals ions in an acid solution to a small volume EMS™ "concentrate" (typically 25% of the feed­stream volume). The EMS™ method coincidentally makes a "clean acid" EMS™ "permeate" (typically 75% of the feed­stream volume). In general, EMSTM processing enables 75% of the bleedstream volume to be recycled; allows up to 2/3 of the bleedstream acid to be recovered as a 10% solution; and delivers dissolved metals in a highly concentrated stream suited to a. metals recovery or b. reduced cost lime neutralization for a given metals precip­itation relative to full bleedstream neutralization (typically a 30% lime use savings). Implicit in the recovery of process water for re-use and reduced reagent addition requirement for metals precipitation is a significant reduction in wastewater-sludge dis­posal facility size.

Jan 1, 1998

By R. L. Loofbourow

In addition, dewatering is seldom complete because of aquifer irregularities and an operator's natural reluctance to start a costly study program, provision of wells, and pumping long before mining begins. In some cases, water pumped from the mine is used in treatment. This use is suggested as a way to minimize effluent effects. The true objective of water supply, however, is to provide water of satisfactory quality at the rate needed and at the least cost. Pumping to permit efficient mining isn't likely to coincide strictly with that objective during much of the mine life. The history of mining is replete with efforts, some highly ingenious, to control mine water. Methods that might be used to tackle the challenge Operators of wet mines are faced with both obvious pumping costs and some unquantified water costs. It is possible to reduce both types of costs, generally by reducing inflow to the mine area-and a surprising number of methods have been used to do this. Both the obvious and unquantified costs at a wet mine are now multiplied by: •Increasing costs of power, labor, equipment, and supplies; •Work at greater depth and, in some cases, in wetter ground; •Increasing use of hydraulic fill; and •Increasing zeal in protection of surface and groundwater quality. There are many easily overlooked ways in which mine water may add to hazards and costs (Loofbourow, 1973). As to their importance, consider these examples: •Increased maintenance of equipment, tires, roads, and ditches; •Increased difficulty in supplying proper ventilation; •Increased hazards from ice and use of electric power; • Loss of friable ore minerals and fines stuck on conveyances and in chutes; •Need to use more expensive methods and explosives; and •Decrease productivity of equipment and personnel. Expressed as cost per unit of ore produced, these unquantified costs can be high. Ideally, advanced dewatering should eliminate most of them. Advanced dewatering is a highly effective water control method, but it isn't problem free. Mines in ideal ground for advanced dewatering are most unusual. To be effective, water must be pumped at a comparatively high rate for a period longer than the life of the mine, from below mining depth. Effects on both groundwater and surface water are likely to be drastic. of increasing mine water costs are listed in the accompanying table. Ten groups of possible methods are cited. Most have been demonstrated in mining; two are known from industries other than mining.

Jan 12, 1980

By Dapeng JIANG, Haixiao YANG, Shanxi WU, Jinwang Zhang, Guoliang XU, Yipeng ZHANG

The subsidence and deformation characteristics of strata with thick unconsolidated overlying strata is quite different from that with consolidated overlying strata. In this paper, based on the engineering background of large mining height mining in No.21 coal seam of Zhaogu No.2 Coal Mine in Henan Province, China, the PFC 2D numerical simulation and surface subsidence field observation was conducted to study the subsidence characteristics of strata with thick unconsolidated overlying strata (680m). Numerical model with dimensions of 2000m 750m was established and mining process with different working face lengths (150 m,200 m,250 m,300 m) and different working face advancing distances (40 m,80 m,120 m,160 m,200 m,240 m,280 m,320 m,360 m,400 m,440 m,480 m) were simulated. The surface subsidence characteristics and displacement field variation of rock strata under different working face advancing distances were observed and analyzed. The results show that with the advance of working face, the structure of subsidence-like basin will be formed in each rock stratum, and the influence range of mining on rock stratum is “bowl-shaped”, and the influence increases with the increase of advancing distance. The unconsolidated layer has a buffer effect on the mining-induced subsidence, and the buffer effect is realized by the expansion of the subsidence area. The movement law of strata under different working face lengths is similar to that under general mining, and the structure of subsidence-like basins will also be formed in each stratum. The influence range of strata after mining is ' bowl-shaped ', and the influence extent increases with the increasing of working face length; the internal subsidence of the unconsolidated layer is more intense and complex than that of the rock layer; the surface subsidence movement shows an obvious 'basin' phenomenon. The field observation data of surface subsidence also confirm the above results.

Compared to alternative ways of dealing with coal fines, oil agglomeration has always been a commercial bridesmaid and never yet a bride. Yet interest is high in the process, as shown by a number of pilot plants in operation that are funded by industry as well as government. In the past in the US, alternative ways to handle coal finer than 100 to 200 mesh included flotation, or no treatment except dewatering, or discarding the fines as waste. There are now a number of factors that tilt the economics more toward fine coal recovery. Today, the cost of mining a ton of underground coal runs from $22-30, and discarding even 10% as unrecovered fine coal is a real economic loss. A number of attempts (notably by EPRI) have been made to quantify the effect of coal cleaning on utility boiler operations. Transportation and ash handling costs, frequency of unscheduled outages, fouling and slagging problems, and boiler tube erosion are all lowered by cleaning coal. The advantage of coal cleaning on these parameters has been shown, but the dollar amount varies with many assumptions and is difficult to pin down. The message to a utility, however, has not been lost and the trend toward cleaner steam coal is accelerating. This may lead to fewer installations that only wash the coarse coal, and add back the unwashed fine coal until a maximum contractual ash limit is reached. And, as the proportion of coal fines continues to increase by continuous and longwall mining, it may become impossible to clean the coal sufficiently without cleaning the fines. Flotation and oil agglomeration are the only processes that presently deal directly with cleaning the very finest coal fractions. Flotation has proven its effectiveness in cleaning blackwater and in separating fine coal from mineral matter. It has the achievable potential for far greater pyrite rejection than is now obtained in commercial preparation plants. While flotation will continue to be the predominant method for fine coal cleaning, there are conditions and opportunities where oil agglomeration can be a

Jan 8, 1980

By Dianne M. Satalic, Peter A. Spencer

Bacterial leaching is a mature technology and has been applied in six commercial plants with others in the design stage or under construction. Bacterial oxidation is a relatively inexpensive process which uses naturally occurring soil bacteria to assist in extraction of base and precious metals. Bacterial oxidation has ini­tially been applied for gold recovery but the technology is now being transferred to base metal leaching with processes already developed for copper, cobalt and nickel. Bacterial oxidation plants often have capital and operating cost advantages over competing technologies and do not need a highly skilled workforce. The most common species found in the conditions used for bacterial oxidation and bacterial leaching are Thiobacillus fer­rooxidans and Thiobacillus thiooxidans. One common species other than Thiobaci!lus found in the commercial bacterial leach cultures is Leptospirillum ferrooxidans. Mixed cultures of the three species named here have formed the basis of the technolo­gy applied in most of the commercial applications of bacterial leaching to date. These bacterial species already mentioned all belong in the classification of mesophiles and have an operating range from 20 to 43°C. One other culture, a moderate thermophile operating over the range 40 to 50°C, has been used in one commercial plant in Western Australia. Bacterial oxidation takes place in acidic solutions under aer­obic conditions. The oxidation reaction normally produces iron (II) in solution which is ultimately conve1ted into the iron (III) form. The pH value falls naturally as the sulphide minerals are oxidised and the starting pH value can fall as low as pH 0.6 as the oxidation reaction proceeds. This paper examines the impact of the technology selection and its application on the design, engineering and operation of a bacterial oxidation plant. The effect of the changes to these para­meters on the cost of the bacterial oxidation plant is discussed.

Jan 1, 1998

By J. J. K. Daemen, C. Fairhurst

Two aspects of tunnel support design are considered, and for both particular attention is paid to the influence of rock failure around the tunnel on the loads developed on the supports. 1) In general it has not been possible to extend analytical methods for the study of rock failure around tunnels to three dimensions. As a consequence it was not possible to calculate either the influence of the face or the actual sequence of support loading as the face advanced. The finite element method made this extension possible. A simplified analysis of the influence of the face and of the sequence of support installation has been made. The influence of the original stress field and of the rock behaviour on the length of tunnel which can be left unsupported (related to the stand-up time) is shown. The degree to which a 2-dimensional analysis can be used to approximate the final support load is shown to be dependent upon the post-failure behaviour of the rock. 2) Tunnel support design requires a knowledge of the forces and of the -rock displacements which cause stressing and deformation of the support system. A 2-dimensional model incorporating a detailed representation of several support systems, of the rock behaviour in the broken zone and of the interaction between rock and support is given. The model allows the calculation of the stress distribution in the support. The significance of the relative stiffness of support and rock on the stresses developed in the support is shown.

Jan 1, 1997

By Heflin. L.

The Washington Metropolitan Area Transit Authority, WMATA, has constructed over 92 revenue miles of rapid transit facilities including some 14 miles of rock tunnel and 12 miles of earth tunnel. The system map is shown in Figure 1. During construction it is not unusual for contractor to encounter ground conditions resulting in major problems and extra cost. Problems such as major rockfalls and a "stuck" boring machine in squeezing rock are not unusual. In recent years, faced with additional tunnel construction projects, the Authority has sought newer tunneling methods which might be less costly. One new method which has been tried is NATM, the New Austrian Tunnel¬ing Method. NATM, which is applicable to both soft ground, mixed face and rock tunneling, relies on miners and comparatively small machines to excavate in conventional ways, but it is unique in its methods of obtaining "stand-up" time at the excavated face and its method of using shotcrete as an initial lining, frequently with a drainage system to reduce hydro-static pressure on the final lining. Lower costs result from use of a cast-in-place concrete liner, and from a reduction in the number of work operations at the mine heading which provides better daily production. WMATA has recently completed three twin bore tunnels constructed by NATM technology. The first was the Wheaton Tunnel and Section, followed by the tail track tunnel to total some 4200 linear feet cut through jointed rock up to 270 feet below ground surface, and the last was in soft ground at Fort Totten. A

Jan 1, 1993

By R. J. Pindred

In his paper, Davis presents an overview of risk. He also introduces the Capital Asset Processing Model (CAPM) as a foundation for selecting the appropriate discount rate for a mining project. While applying portfolio theory is more defensible than the ad hoc adjustment of discount rates, the CAPM is not a panacea.

Jan 1, 1995

By D. A. Rice

As part of a cooperative effort between the U.S. Bureau of Mines and the Florida Institute of Phosphate Research (FIPR), with input-from the phosphate industry, the conversion of phosphor-gypsum to sulfur is being investigated. The proposed process incorporates a thermal reduction of phosphogypsum to calcium sulfide (CaS) and a hydrometallurgical treatment to convert calcium sulfide to ammonium bisulfide (NH4SH), which is then oxidized using a catalyst to produce elemental sulfur. The paper summarizes the results of continuing laboratory studies using an activated carbon catalyst which was loaded with sulfur through oxidation of ammonium bisulfide solution. The extraction of the sulfur that forms in situ on the catalyst is a key area in the development of the sulfur recovery process. Two approaches for removing the sulfur are being investigated: 1) thermal treatment to volatilize sulfur, and 2) leaching the sulfur using liquid anhydrous ammonia. The research has shown that it is technically feasible to recover over 90% of the sulfur from the carbon by either volatilization or ammonia leaching. Locked-cycle testing has demonstrated that the carbon has the potential to be recycled after ammonia leaching.

Jan 1, 1988

By Ki Chun

Optimizing ore processing requires a thorough understanding of not only the process itself but the economic constraints which influence business decisions. No single modeling technique is able to model all the different processes running in today's mines. The solution then is to use a number of different model­ing techniques, each competing with the others to best correlate the data. Among these modeling techniques are: Neural Networks, Genetic Algorithms, Fuzzy Logic, Crisp Logic, Statistics, First Principal Models, and System Identification models. Once a model has captured the behavior of the process, a numerical optimization determines the optimal operating point and set point parameters. A single step is then taken in the direc­tion of the optimal location by changing all the process parame­ters to the setting recommended by the optimization system. As the process begins to re-stabilize, the models are updated, the new optimal determined and another step is taken in the direction of the optimum. Through the use of multiple competing models integrated into a numerical optimization the process can achieve the optimal running condition even if the ambient conditions fluctuate dynamically.

Jan 1, 1998

By Sean E. Armstrong, Deepak Malhotra

Gold extraction from refractory ores as a percentage of total gold production is increasing. The processing technology options for treating difficult-to-process refractory ores are also increasing. The best processing option is dependent on many factors including feed grade and the cause for ore refractoriness. The causes for ore refractoriness are many foldCpreg-rob­bing, encapsulation in pyrite, arsenopyrite or silica, presence of cyanocides, presence of cyanide-insoluble gold compounds, etc.1. The first step in the process development of a refractory gold deposit is to determine the cause for ore refractoriness. The traditional approach of undertaking detailed mineralogical examination of ore types, major and minor mineral constituents and their association does not necessarily provide the metallur­gist all the information needed for characterization of gold-bear­ing ores. For process development programs it is desirable to complement traditional mineralogical methods with diagnostic leach testing, which attempts to quantify the deportation of gold in various mineral matrices. This paper briefly reviews the diagnostic leaching proce­dures and its applications for a wide range of gold-bearing refractory ores, both primary and by-product.

Jan 1, 1996

By Darryl J. Butcher

The Akjoujt Project is operated by MORAK SA, a company incorporated in the Islamic Republic of Mauritania. MORAK is jointly owned by General Gold International SA (A wholly owned subsidiary of General Gold Resources NL.), SAMIN, and the International Finance Corporation. The project was devel­oped to treat approximately 2.1 million tonnes of TORCO (Treatment of Refractory Copper Ores) tailings containing an average 3.2gAu/t, and 0.75%Cu/t. The process involves the use of ammoniacal cyanidation for gold extraction, and conventional carbon adsorption, and eventual desorption, of gold values. This is believed to be only the second commercial application of this leach chemistry in the world. The other being a small vat leach operation on the Paris Mine Dumps. (This operation was described by Mr M Ruane in a paper to a previous Randol Forum.) Initial work on the process, and the Akjoujt process design assumed that the reaction involved was anaerobic. This paper describes the operating experience, plant trial, and eventual installation which has lead MORAK to the conclusion that the process is not anaerobic, but requires oxygen, or an oxidant to proceed.

Jan 1, 1995