energy efficiency energy intensity in copper and gold mining
- mineral processing
Summary: Mines are faced with numerous challenges, such as falling raw material prices, declining metal grades in the ores and higher energy prices. Especially because ore processing is particularly energy-intensive, the industry is again focusing on the saving of energy. This report shows what are the key energy saving considerations in the copper and gold mining sectors.
In recent years, many mining companies have been able to reduce their specific energy requirements in the ore processing of base metals, gold, silver and platinum group metals, thereby improving their competitiveness. The reasons are complex and involve, for example, the closure of unprofitable mines, technological improvements such as modern grinding processes or improved energy management. But there are also opposing tendencies. For example, the electricity demand for Chiles copper production is expected to increase by 53.5% between 2015 and 2026, although the planned increase in copper production over that period is only 7.5%. An analysis of the reasons for this reveals that the contributing factors are not only the type of ore dressing process, but also the declining ore head grades and the supply of water to the mines.
After iron ore, the leading places in the global sales ranking for mineral raw materials are occupied by copper and gold. Despite an interim sales crisis, the demand for copper and gold is unbroken. Fig.1 shows the development of copper mine production over the last 10 years. Production has risen at average annual growth rates (CAGR) of 3.0% to 20.2milliont (Mt). The annual growth fluctuations were between 8.9% and -0.2%. In the case of gold production, the situation is similar (Fig.2). The CAGR of gold production was 3.6% in the last 10years. In 2016, 3260tonnes of gold were produced, after 2350tonnes in 2007 and a relatively small slump in gold production during the financial crisis of 2008.
In principle, it is clear that higher productions can only be realized with a greater energy expenditure. This technical paper will particularly consider the electrical energy input in more detail. The electrical energy input and the fuel requirements are roughly the same in the gold and copper production sectors. Of particular interest is the specific energy input per ton of recovered copper or gold. The relevant data are provided primarily by the mining companies. To a lesser extent, data are available from individual mining associations. Further data sources are energy audits and the project descriptions of consulting companies for mine expansions and new copper and gold mining projects. The situation is fundamentally similar in the case of other mineral raw materials.
At present, copper production is stagnating in Chile (Fig.3). In 2013, a peak mining output of 5,776Mt was reached, while in 2016 the production figure was 5.553Mt, corresponding to a share of 27.5% in the worldwide mining output of 20.2Mt. Starting from 2015, the Chilean Copper Commission (Cochilco) is planning to increase the annual mining output by 0.5% in order to achieve about 6.2Mt by 2026. In 2015, Cochilco published a study on the future electricity demand of the Chilean copper mining industry. This study covers all current and future projects. Fig.4 shows how the future electricity demand is forecast to change from 22.1terawatt hours (TWh) in 2015 to 34.1TWh. The various consumers are also shown.
The largest consumer is the conventional ore dressing process with a concentrator as used for sulfidic ores, consisting of crushing, grinding and subsequent flotation. This is followed by processes for oxidic copper ores with heap leaching, solvent extraction and electrowinning (LS-SX-EW). The share of ore production by concentrator processes will increase from 72% in 2015 to 89% in 2026. Correspondingly, the electricity demand for the concentrator processes will increase by 69% from 13.2TWh to 22.3TWh in 2026 while that of the LS-SX-EW processes will decrease by 40% from 4.5TWh to 2.7TWh. Also interesting is the increase of almost 460% for seawater desalination/pumping. The increases in mining (+38%), refining (+30%) and services (+26%) within the framework of the planned additional mining output are somewhat more moderate, but clearly exceed the planned production growth.
For many years the copper ore grades have been declining as exploitation of the higher-grade deposits progresses. On a worldwide scale, the mined ores contain an average of less than 1.0%Cu (Fig.5). In Chile, the copper grade of the ores is significantly lower than even that. In 2015, the average copper grade in the ores was only 0.65%. For the year 2026, it is expected that the ores in Chile will have copper grades of less than 0.5%. This will naturally increase the amount of run-of-mine ore that has to be processed. In international comparison, the Chilean copper ore mining industry is steadily deteriorating. While about 35% of the worldwide copper mine output had higher copper grades than the product of Chilean mines in 2010, this figure will rise to 43% in 2020.
Fig.6 shows the development of the electricity demand for concentrator and LS-SX-EW processes. While in the case of concentrator processes the specific electricity demand per t of material has increased by 4% from 79.3MJ/t to 82.5MJ/t, the power demand for the LS-SX-EW processes has decreased by almost 23% from 43.2 MJ/t to 33.3 MJ/t. The figure also shows that the power demand per t of material is significantly lower in the case of LS-SX-EW processes. However, in the concentrator processes this disadvantage is compensated by higher yields. As depicted in Fig.7, the specific electricity demand per ton of produced copper is currently about 12.0GJ/t of copper (Cu) in both the concentrator and the LS-SX-EW processes. However, in the case of plants equipped with concentrators, this value has increased at a faster overall rate over the past 10 years.
The mineralogical properties of the ores are increasingly influencing the required processing methods. In the production of copper, sulfidic ores are enriched into concentrates by grinding and flotation, followed by pyrometallurgical processes for the production of pure copper. Oxidic ores are treated with sulfuric acid in a heap leaching process after the grinding stage and are subsequently processed into cathode copper by SX/EW methods. On a worldwide scale, concentrate production dominates with a market share of about 85%. In the production of gold, the cyanide leaching or CIL (carbon-in-leach) processes have gained a market share of almost 90%. For the worldwide energy demand of these processes only estimated figures are currently available.
Fig.8 presents a simplified overview of the energy input for the different processes and process stages in the production of copper based on a copper ore grade of 0.5%. The energy inputs are expressed in kJ/t of material and kJ/lb (pound of copper). The so-called run-of-mine (ROM) leaching followed by a SX/EW process has the lowest energy demand, while the process with SAG/ball mills and subsequent flotation and pyrometallurgy has the highest energy demand. For each respective process, various possibilities for reducing the energy input are shown. Fig.9 additionally shows how the energy demand for the different processes varies depending on the copper grade of the ore. Correspondingly, the respective theoretical energy inputs range from less than 10MJ/lbCu up to more than 80MJ/lbCu.
The greatest energy input in copper and gold production is required for the comminution and grinding processes. The energy audit of mainly Australian copper and gold mines shows that 36% of the overall energy consumption is attributable to comminution . Previous studies had shown values between 18% and 50%. Fig.10 shows that the specific comminution energy is a function of the copper grade of the ore, but also a function of the throughput of the mine and thus of the technology employed. Therefore, on the one hand, the declining copper grades require higher specific energies of more than 4MWh/tCu, while on the other hand, so-called scale effects occur in the case of larger mines and partially compensate for poorer copper grades, making specific comminution energies of less than 1MWh/tCu possible.
The classical grinding process employing SAG and ball mills (Fig.11) is progressively losing ground [2,3] against high-pressure grinding rolls (HPGR), which are increasingly being used for the grinding of both copper and gold ore. The first noteworthy HPGR application was at the Cyprus Sierrita copper mine in the USA in 1995, even though the extremely abrasive ore prevented the achievement of economic grinding roll service lives. This changed with the Cerro Verde project, a copper-molybdenum mine in Peru in 2006 and later in 2011, when the SAG mills were completely replaced by 4HPGRs, each with a capacity of 2100t/h. In 2014, the largest HPGR that has been implemented so far was installed in Freeport-McMoRans Morenci copper mine in Arizona/USA, achieving throughputs of up to 5400t/h (Fig.12). Pilot studies had demonstrated that the HPGR can achieve energy savings of 13.5% compared to grinding processes employing SAG mills. By 2014, more than 35HPGR had already been put into operation in the copper grinding industry.
A further important area for power consumption reductions is the after-grinding of products from the flotation stage [3, 6, 7]. Here, the fineness requirements are in the broad range of 2 to 75m, the feed particle sizes are smaller than 200m and the throughput rates are usually below 100 t/h, so that due to their high energy requirements ball mills do not represent a reasonable solution in this range of fineness. With the IsaMill horizontal stirred media mill (Fig.13), it is possible to achieve energy savings of 20-30% compared to conventional ball mills in various applications. High energy savings are also achieved with vertical flow stirred media mills (Fig.14). Due to their relatively low energy requirements, both these mill types are also being increasingly used for secondary and tertiary grinding as replacements for ball mills [8, 9, 10].
The decreasing mineral grades in the ores have also led to increased throughput rates in the flotation stage (Fig.15). In order to accommodate higher flotation volumes, plant manufacturers have undertaken a scale-up of the flotation cells. So-called supercells are now available with volumes of up to 600m3. It is of particular interest in this context that the larger cells provide an improved hydrodynamic performance compared to small cells using conventional technology, a factor which reduces energy costs by up to 40%. However, the amount of savings thus achieved for the entire processing line must be put into perspective, since the flotation stage usually accounts for less than 10% of the energy costs for the grinding stage.
Two other areas that are also among the notable energy consumers are the mechanical handling of the material and the pump conveyance of liquids and slurries. Mechanical conveyor systems (Fig.16) are indispensable for feeding the processing equipment. In addition, there are various mechanical conveyors in the mine itself. The consumed electrical energy amounts to less than 5% of that of the total ore processing line. The energy consumption situation for pumping is somewhat different. In particular, when seawater desalination plants and long-distance water supply are included, the electrical energy consumption can rise to over 15% of the total. High energy losses occur, in particular, when slurry pumps suffer premature wear. This problem can be reduced by special designs, such as wear rings in the pump.
Australian scientists have evaluated the energy consumption of a total of 68 copper and gold mines covering all such mines in Australia as well as 24% of the worlds copper mines and 15% of the worlds gold mines. While complete data for SAG and ball mills were available, only partial data were obtained for crushers and post-grinding, so that some results were estimated. Fig.17 presents the results for the specific comminution energy for the grinding of copper ore. The average consumption is 1.223MWh/t Cu. On the abscissa, the cumulative copper production quantity of the analyzed mines is shown. A corresponding graph also exists for the analyzed gold mines (Fig.18). This represents mines with a production of 11million ounces(Moz). The average specific comminution energy is 353kWh/oz.
Fig.19 depicts the specific energy consumption of the companies Teck Resources and Barrick Gold in the copper ore processing sector. The specific data are shown in GJ/tCu and thus provide a measure of the energy intensity of the copper production process. In the case of Teck Resources, the data represent 4 mines in Canada, Chile and Peru with a total production volume of 324kt in 2016. Barrick Gold has only the Lumwana copper mine in Zambia with a most recent production volume of 123kt. With both companies there is a trend towards lower energy intensities over the past three years. The energy intensity of 24.5GJ/tCu at the Lumwana mine is significantly lower than Tecks average value of 43.7GJ/tCu in 2016.
Fig.20 shows the energy intensity for the company Gold Fields. Gold Fields operates 3gold mines in South Africa, Australia and Ghana as well as a copper/gold mine in Peru. In 2016, the company produced 2,15Moz of gold (calculated as gold equivalent), which was almost equal to the previous years figure of 2.16Moz. For the 2016 gold production quantity an energy input of 0.063GJ/t of ore material was necessary, after the figure of 0.072GJ/t in 2014. The energy intensity for gold production was 5.27GJ/oz in 2016 after 4.56GJ/oz in 2014. This means that the energy intensity increased over the two years with a CAGR of 7.5%. Such an increase can only be explained by a sharp decline in the grade of gold in the ore.
Fig.21 shows the energy intensity of the company Barrick Gold for the gold production of selected mines. In total, Barrick Gold owns 9gold mines in Argentina (Veladero), Canada, the Dominican Republic, Peru and the USA (Cortez, Goldstrike, Turquoise Ridge and Golden Sunlight). In 2016, the company produced a total of 5.52Moz of gold, making Barrick the no. 1 worldwide, ahead of Newmont Mining and AngloGold Ashanti. The average energy intensity of the 9 goldmines decreased from 5.33GJ/oz in 2014 to 5.11GK/oz in 2016. However, the 3depicted mines have significantly different energy intensity levels and graph bar sequences. Goldstrike is among the Barrick Gold mines with the highest energy intensity, although its energy demand has been reduced. Turquoise Ridge has the lowest energy demand with minor fluctuations in recent years. Veladoros energy demand showed an upward trend.
Lower metal grades in the ores force mine operators to look for solutions in order to further reduce the energy demand of their processing lines. If no efforts were made, the energy demand would rise significantly. Mining companies are increasingly carrying out energy audits in order to identify the largest energy consumers and to see how plant performance can be improved. The main focus is on the grinding process, as it is the largest energy consumer. For new projects, more energy-efficient grinding processes should be considered, employing HPGR instead of SAG mills and vertical and horizontal stirred media mills instead of ball mills. An important role in energy demand reduction is played by the optimization of grinding circuits.
In the case of performance enhancements and Brownfield projects, the focus is on adding machines to existing circuits or replacing machines with higher-performance equipment. HPGR mills and stirred media mills also play an important role here as a simple means of achieving higher throughputs with an additional grinding circuit or other machine combinations. This also concerns the post-grinding after flotation and the replacement of existing flotation cells with new, higher-performance and larger cells in order to achieve longer residence times and better yields. In the recent past, ore selection by sensor-based processes has also been intensively discussed. In any case, mine operators today have numerous options for energy saving while simultaneously reducing cash costs.
The top ore layer of an open pit copper mine is easily processed using heap leach in tandem with solvent extraction and electrowinning to produce copper cathodes. The copper mineral most predominate...
ABB was recently contacted to supply the drive system for a new semi-autogenous grinding (SAG) mill at Neves Corvo/Portugal. The contract was awarded in June 2011. Neves Corvo is an underground mine,...
cyanide-free gold processing technology hits the market
The process replaces
cyanide with a safer, less hazardous chemical reagent called thiosulphate. This
inorganic compound helps dissolve fine gold out of ores into a solution, which
can then be recovered through further processing.
The new technology
was developed over more than a decade by Australias national science agency,
CSIRO, and trialed in Australia in 2018 with Clean Minings parent company
Eco Minerals Research Limited. This trial proved the thiosulphate solution
could extract gold from ore at an industrial scale.
and the associated tailing dams from the gold recovery process is a
game-changer for the sector and, importantly, for the communities where gold
miners operate, Mr McCulloch said.
Clean Mining will
initially target small to mid-scale miners who can benefit from the
cost-effective leaching ore processing solution, which includes a plug-and-play
plant that can be customized and scaled to meet individual needs.
Would you please send more information about thiosulphate.as more for small scale gold mining . Part of family working in small scale gold mining. This section employs more e than 1 million. Please give some information about the price. and the possibility to be you execlutive agent in Sudan,
red lake gold mines - canadian mining journal
Rivals since 1949 when they began pouring gold, the Campbell and the Red Lake mines have at different times each been called the worlds richest gold mine. Now a single owner, Goldcorp Inc., is combining the adjacent operations in Balmertown, Ont., reducing costs and nearing its 1-million-oz/year target.
Combining two mines, mills, cultures and workforces that have fostered a rivalry for over 50 years does not happen automatically. Red Lake Gold Mines (RLGM) mine general manager Dan Gagnon said the cultures are melding. We are slowly growing together. This is a small town of 4,600 people and they see each other at the rink and the bank, so they talk about what is happening and this helps with integration.
Everything at Campbell and Red Lake used to be done separately and differently. Weve had to look at everything and choose the best options to go forward, and this by taking the best of both worlds. The two mines and mills had differing wage and benefit packages, differing approaches to labour (contract vs. full-time employees), and in the warehouse commonality extended to only 11% of the parts and supplies.
Now, as the Red Lake complex and Campbell complex come together under the same management, plans call for gold output to be 786,000 oz in 2007 (70% from Red Lake and 30% from Campbell). By 2011 that number will reach 1.0 million oz, thanks to continuing production efficiencies and the superior exploration potential of the Red Lake camp.
Goldcorp has drawn up a US$26-million budget for exploration this year in the area of RLGM. With that investment the company plans to keep 15 rigs busy with 180,000 m of drilling and complete 610 m of underground exploration development. Targets have been chosen in five key areas.
The HGZ is a particular priority because grades are going up as drilling goes deeper. The HGZ above 37 level averages 51.4 g/t (1.5 opt) Au, and between 37 and 47 level averages 115.9 g/t (3.4 opt) Au. Proven and probable reserves at the end of 2006 to the 47 level totalled 1.4 million tonnes grading as high as 99.5 g/t Au. In terms of contained ounces, the HGZ has 3.6 million oz in reserves. Drilling has outlined resources containing 1.4 million oz from 47 to 49 level. And the mineralized zone is known to extend below 51 level.
Parallel to the HGZ runs the footwall sulphide zone interpreted as a group of five separate zones. The sulphides have a geometry ideal for longhole development. Their location between 27 and 37 level will reduce reliance on extremely deep ore sources as mining reaches these zones in 2008 and 2009. Sulphide reserves identified below 30 level total 969,000 t with an average grade of 8.4 g/t Au. Drilling continues from the 34 level.
The DC zone lies below the 41 level of the Campbell mine. Not only are high gold values being intersected, but the zone is very continuous and open both below 50 level and to the west. Reserves have been calculated in the DC and DCE zones as 390,000 t grading 19.9 g/t Au (0.6 opt) Au, and resources in the PCB and TP zones are 3.7 million t grading 9.9 g/t Au. Between the 41 and 48 levels, these structures may contain over 1.2 million oz of gold. Exploration drilling continues from the 4199 hanging wall exploration drift.
Combining the Red Lake and Campbell mines under one owner has the obvious advantage of allowing development of the Party Wall area. This is the undeveloped no-mans-land between the two mines when mining ceased at the property boundaries of the respective owners. The Party Wall will no longer be a data void; Goldcorp is outlining resources in the area in preparation for mining from the Campbell workings.
Drill-testing the Party Wall from 19 to 25 level has outlined a resource of 204,000 t grading 13.4 g/t (0.39 opt) Au in the E and F zones. The drilling program will now test the west extension and down dip to 29 level.
The RLC sulphides, last mined in 1966, lie near the surface and are accessible from the Campbell workings. By reinterpreting the geological model, Goldcorps geologists believe the multiple zones include a million ounces of minable gold. From surface to 30 level the zones contain reserves of 620,000 t grading 15.1 g/t (0.44 opt) Au and resources of 1.6 million t at the same grade. Plans are being made to mine these sulphides again, as a means of reducing dependency on extremely deep ore resources. Mining is scheduled to begin in 2008 using mostly longhole stoping methods.
Various surface targets have been chosen between the Reid shaft at the Campbell mine and the No.1 shaft at the Red Lake mine. A previously unknown vein has been drilled and a test pit was worked last summer. The open pit concept will be aggressively tested in 2007, and if a new pit is to be built at the RLGM, this would be a first in the area.
The Red Lake gold mine was the first Goldcorp producer. It operated as the poor sister to the extremely high-grade Campbell mine until Goldcorp developed the HGZ at depth in preparation for reopening after a long labour stoppage. And high-grade is the name for reserves topping 3.0-opt Au.
The developed area includes a production shaft (No.1) to 1,023 m (1 to 23 level) and an internal winze (No.2) from 23 level to 38 level. No.3 shaft reached its planned depth of 1,924 m in January 2007. The new shaft will provide more hoisting and ventilation capacity, hence increasing the mining rate from 2,350 to 3,100 t/d, of which 40% is ore. A main access ramp connects workings from 21 level to 42 level, and the ramp will eventually reach 50 level.
Mining is mostly cut-and-fill, both overhand (37%) and underhand (40%). Other ore sources are pillar recovery (8%) and longhole stopes (15%). Muck is hauled to ore passes on 34 and 37 level with load-haul-dumpers (LHD) ranging in size from 0.75 to 2.7 m3. Ore is hauled on 38 level by train to the bottom of No.2 shaft, hoisted to 23 level, and skipped to the surface using either of two 5.5-t skips in No.1 shaft.
The Campbell gold mine is a recent addition to Goldcorps list of producers. Goldcorp bought the mine in 2006 from Barrick Gold as part of the Barrick takeover of Placer Dome and its subsequent sell-off of selected assets.
The Campbell shaft has been deepened four times, reaching 1,316 m and servicing 27 levels. In 1999 construction began on the Reid shaft 150 m to the west. It was sunk to 1,819 m and opened up deep ore sources including the DC zone. A ramp has been driven from the 39 level at the bottom of the Reid shaft to the 45 level and provides a drill drift from which the ultimate depth of the DC zone is being tested.
A variety of mining methods are used in the Campbell mine. Down to the 27 level, a combination of mechanized diesel equipment and tracked haulage is practised on every level. Below 27 level all mining is trackless. This year plans call for mining 532,000 tonnes of ore from longhole (65%), cut-and-fill (30%) and development (5%) stopes. Ore is hoisted through the Reid shaft.
The original Red Lake mill was built in 1948 and was replaced by a new facility in 2000. Throughput at the new mill is ramping up to 1,250 t/d by the end of July 2007. The mill consists of three distinct parts: two-stage crushing; gravity and CIP circuits followed by electrowinning or autoclaving; and a paste backfill plant.
The gravity circuit (a ball
mill, two Knelson concentrators and Diester table) recovers about 55% of the gold, making a 75%-Au concentrate that is directly smelted. The CIP circuit recovers approximately 34% of the gold. Slurry is leached in four tanks, passes through the CIP circuit, and the pregnant solution reports to a pair of electrowinning cells. The precipitated gold-bearing sludge is directly smelted.
The Red Lake mill also produces a bulk sulphide concentrate using a conventional flotation circuit. It is shipped to the Campbell mill to be autoclaved. About 9% of the gold is recovered from this portion of the ore.
The Campbell mill treats both free milling and refractory gold ore at a rate of 1,850 t/d. Run-of-mine ore is crushed in three stages to 19 mm and then enters a two-stage rod and ball mill circuit. The slurry is pumped to a bank of eight cyclones, the overflow going to the flotation circuit and the underflow going to the ball mill. A pair of Knelson concentrators and shaking tables recover approximately 40% of the gold by gravity means.
The cyclone overflow reports to a bank of seven Denver DR-500 rougher cells, and the resulting concentrate enters a bank of four Denver DR-100 cleaner cells. Cleaner tails are recycled back to the roughers at the head of the circuit. Flotation tails are thickened, with the underflow being sent to the leaching circuit and the overflow to the process water tank.
Thickened flotation concentrate feeds the pressure oxidation circuit. Pressure oxidation replaced the antiquated roaster circuit in 1991. After pre-treatment, the slurry flows by gravity through five chambers in the autoclave. This is a 2.8-m-long by 12.2-m-diameter vessel that operates at 2,100 kPa and 190C sustained by the exothermic nature of the chemical reactions. Designed slurry retention time is two hours.
Depressurized and slightly cooled slurry leaving the autoclave passes through a two-stage CCD wash circuit. The CCD second wash underflow is transferred to the carbon-in-leach (CIL) circuit. The oxidized slurry undergoes cyanidation and carbon adsorption in a pair of tanks, each having a retention time of 48 hours.
The leached concentrate from the second CIL tank is combined with thickened flotation tails and further treated in a carbon-in-pulp (CIP) circuit. It consists of six tanks, each having a slurry retention time of 50 minutes. After the carbon is stripped, the pregnant solution enters the electrowinning cell.
With the one-million-oz/year target clearly within range, Goldcorps RLGM is exhibiting the best of both worlds: the high-grade (45.0 g/t Au), low-tonnage Red Lake mine and the high-tonnage, lower-grade (12.0-15.0 g/t Au) Campbell mine. The company is fortunate to have two proven producers and a host of exploration potential in one of Canadas richest gold camps.
gold ore carbon-in-leaching (cil) processing technology - sbm mining and construction machinery
SBMwork together with the famous laboratory in China to design and deliver the gold leaching plant. It help the mine factory to extract a major part of the residual gold contained in gold ore or some tailings. Here is the brief introduction of this very popular gold ore processing technology in the world. We do beleive that your gold ore processing business will be more sucessful by choosingSBM.
The ore is ground and concentrates are produced by means of conventional otation and gravity circuits. The otation tailings containing the unrecovered gold from the primary circuits are directed to the leaching plant and dissolved in an aerated sodium cyanide solution. The solubilized gold is simultaneously adsorbed onto coarse granules of activated carbon directly in the pulp in the so-called Carbon-In-Leach process (CIL). The loaded carbon is treated at high temperature to elute the adsorbed gold into the solution once again. The gold-rich eluate is fed into an electrowinning circuit where gold and other metals are plated out onto cathodes of steel wool. The loaded steel wool is pretreated by calcination before it is mixed with uxes and melted. Finally, the melt is poured into a cascade of molds where gold is separated from the slag as gold bullion.
The CIL process utilizes safe and reliable technology and provides xibility on raw materials analysis. It requires low inventory of precious metals. In addition, it ensures short processing time and low operating costs.
The leach circuit comprises eight stainless steel tanks in series, located outdoors. Each tank is equipped with an agitator. The rst tank is a pure leach tank and the following seven are CIL tanks. The otation tailings slurry is pumped to the st tank via a trash screen. Sodium cyanide, slaked lime slurry and air are added into the tanks. The slurry ws from tank to tank by gravity and gold is leached into the solution by cyanide and oxygen. Lime is added to the leach circuit to maintain a high pH of the slurry. This minimizes the formation of hydrogen cyanide.
The solubilized gold is adsorbed onto coarse granules of activated carbon in the tanks at the same time as the slurry ws from the st CIL tank to the last. The carbon is retained in the tanks by using cylindrical, mechanically swept interstage screens, which are submerged in the pulp. The aperture of the interstage screens allows the leached pulp to pass through by gravity. The activated carbon is advanced counter-current to the slurry w by pumping a portion of the slurry upstream with airlifts. A high grade of gold on the carbon and a high gold recovery are obtained by stage-wise adsorption and a counter-current ow. The carbon in the rst CIL tank thus contains the highest grade of gold.
Loaded carbon from the st CIL tank is transferred by a slurry pump to the loaded carbon screen. The slurry is screened, the oversize carbon washed and gravitated to the vertical elution column. Here the carbon is washed with dilute hydrochloric acid to remove for instance soluble salts. The carbon is further rinsed with water to remove the acid before it is soaked with a caustic sodium cyanide solution at elevated temperature. A number of bed volumes of hot water are then pumped through the bed of carbon to eluate the gold. The gold-containing eluate is collected in the electrolyte tank.
Not only gold is adsorbed onto the activated carbon, but a number of organic components, such as oils and otation reagents, are accumulated within the porous structure of the carbon, blocking the active surface. A major part of the organic species remains on the surface after the elution process. The carbon also loses activity during the adsorption of gold and other species. In order to restore the active surface and to enable the reuse of carbon in the CIL circuit, it undergoes thermal regeneration. This is done in a gas-heated horizontal kiln at 650, under a steam atmosphere.
The gold-rich solution from the elution process, the electrolyte, is circulated through three parallel electrowinning cells over a period of 1624 hours. Gold and other metals are precipitated on steel wool cathodes, which are submerged in the circulating electrolyte. At the anodes, oxygen is formed and degradation of cyanide occurs. The barren electrolyte is pumped to the process for reclamation of residual gold and to recover cyanide and sodium hydroxide.
The loaded steel wool cathodes are transferred to an electrically heated calcination furnace. Iron and base metals are oxidized by air at about 750. The obtained calcine is smelted together with xes in a gas-heated furnace. Finally, the melt is poured into a cascade of molds where gold is separated from the slag as dor bullion. The oxidized impurities are present in the slag, which ats on the molten gold. The gold content in the final products is depending mainly on the gold-to-silver ratio in the calcine. Silver follows gold through the process.
The barren pulp from the CIL circuit and any cyanide containing spillage are pumped to the cyanide destruction circuit via a carbon safety screen, to detect any leakage of carbon from the last tank. The barren pulp is combined with the non-cyanide-containing tailings from the parallel otation circuit. The residual cyanide content is destroyed in two reactors by means of the INCO/SO2/Air process before it is pumped to the tailings pond. Cyanide is oxidized to cyanate by adding air, sulphur dioxide and lime in the reactors. Copper sulphate is also added when required. Cyanate and weakly soluble salts of cyanide are precipitated. Cyanate slowly decomposes in the tailings dam to nitrite, nitrate and nitrogen via formation of ammonia/ammonium. Residual-free cyanide and nitrogen compounds in the efent are negligible and do not have any impact on the environment.
Like all the other mineral ore pcoessing plant, the gold ore processing plant should be designed base on the dressing test to fix the dressing index for the whole plant include such as the particle size, pulp density, reagent type and dosage, pickling time, desorption temperature, etc. Here is the datas after dressing test we fixed for this plant. 5). Any pulp collected inside the bunds or in the emergency pond can be pumped back to the CIL or destruction circuit. Bunds are present around all equipment where there are risks of cyanide solution leakage and they make it possible to collect and pump solutions back to the process.
1). Cyanide has to be handled with caution. It is delivered to the plant in bigbags, packed in wooden boxes in sealed steel containers, and stored in a locked area located in the reagent storage building.
2). It is important to maintain a high pH value in the cyanide leaching process to minimize the formation of hydrogen cyanide. The pH is monitored by on-line measurements and controlled by the addition of slaked lime slurry. Too low a pH in the tanks will set off an alarm. The on-line pH probes are calibrated and checked against portable pH meters on a regular basis.
3). Equipment located indoors is covered and well ventilated where there is a risk of a high concentration of hydrogen cyanide. Ventilation gases are cleaned in a wet caustic scrubber system at high pH before they are vented off to the atmosphere. A small volume of cyanide containing scrubber solution is regularly diverted to the CIL circuit and replaced by an equal volume of fresh caustic solution. The ventilation system is also connected to an emergency power supply system.
4). All leaching tanks are located outdoors on a concrete foundation constructed with bunds, which can contain any pulp leakage from the tanks. The bunds also protect the tanks from accidental damage from loaders and other vehicles. The foundation overws to an emergency pond and is heated to keep the or inside the bunds free from ice and snow during the winter period.
5). Any pulp collected inside the bunds or in the emergency pond can be pumped back to the CIL or destruction circuit. Bunds are present around all equipment where there are risks of cyanide solution leakage and they make it possible to collect and pump solutions back to the process.