a new mining concept for extraction metals from deep ore ... · metsim (software) steady-state...

A new mining concept for extraction metals from deep ore deposits by us-

ing biotechnology

D 7.2 CAPEX/OPEX analysis and assessment of the entire metal leaching and recovery

process

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Umwelt- und Ingenieurtechnik GmbH Dresden

With the collaboration of Cobre las Cruces, S.A.

Checked by: Approved by: Name: Dr. Horst Märten, UIT Name: KGHM Date: 2018-07-20 Date: 2018-07-30

Signature: Signature:

Public Document

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D7.2

Due date of Deliverable 2018-07-31

Actual Submission Date 2018-07-31

Start Date of Project 2015-02-01

Duration 42 months

Deliverable Lead Contractor UIT GmbH Dresden

Revision Version 1.0

Last Modifications 2018-07-31

Nature R

Dissemination level PU

Public Summary enclosed no

Reference / Workpackage WP7

Digital File Name De-180831-0052 - D7.2 CAPEX, OPEX analysis and assessment of the entire metal leaching and recovery process

Document reference number De-180831-0052

No of pages 83 (incl. cover and annexes)

Keywords

In bibliography, this report should be cited as follows:

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List of figures

Figure 1: Remediation options and their effect on the post-mining phase .............................................................................................................. 23

Figure 2: Schematic of ISR, in particular, emphasizing the consumption of chemicals, energy and water (C, E, W, respectively). The BIOMOre options (in red circles) include ex-situ biooxidation and the potential in-situ bioleaching (both requiring O2). ...................................................... 24

Figure 3: ISR feasibility criteria, wellfield design and performance and link to the economic model (D7.1) ......................................................... 26

Figure 4: Copper extraction through hydrometallurgical processing ..... 27

Figure 5: Scheme of proposed technology for copper and other metals recovery by hydrometallurgy methods. ................................................. 31

Figure 6: CLC process simplified scheme ............................................. 32

Figure 7: Copper extraction simplified scheme compared to the U ISR process. ................................................................................................. 35

Figure 8: Zinc/Lead recovery routes. ..................................................... 36

Figure 9: Direct leach process with combined ferrite/sphalerite leaching in one stage. .......................................................................................... 37

Figure 10: Combination of Zn current technology and direct leaching .. 37

Figure 11: Operating costs for zinc production dependent on the process time evolution. ....................................................................................... 38

Figure 12: KiLea-Hy leach performance model – metal concentration in PLS and in-situ recovery (left) and daily production and ore resource over the time (right) for the ISR scenario described in the text. ............ 41

Figure 13: CAPEX cost factors for oxidative Cu ISR production scenario described in the text. Total specific CAPEX: 0.48 €/kg. ........................ 42

Figure 14: OPEX cost factors for oxidative Cu ISR production scenario described in the text. Total specific OPEX: 2.81 €/kg............................ 42

Figure 15: As for Figure 14, but assuming non-oxidative Cu ISR (i.e. leaching from oxidized Cu minerals). Total specific OPEX: 1.09 €/kg. .. 43

Figure 16: As for Figure 13, but assuming a deposit depth of 50 m only (instead of 250 m). Total specific CAPEX: 0.14 €/kg............................. 44

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Figure 17: Normalized specific ISR costs (CAPEX + OPEX) in dependence on the calcite abundance in the ore deposit (red line: reference case at 0 wt% calcite) ........................................................... 45

Figure 18: Normalized specific ISR costs (CAPEX + OPEX) in dependence on ore grade (red line: reference case at 1 wt% copper) .. 45

Figure 19: Normalized specific ISR costs (CAPEX + OPEX) in dependence on the depth of the ore body (red line: reference case at 250 m depth) ......................................................................................... 46

Figure 20: Normalized specific ISR costs (CAPEX + OPEX) in dependence on permeability corresponding to about the porosity figures included (red line: reference case 20 % porosity or 0.6 m/d permeability) .............................................................................................................. 47

Figure 21: Normalized specific ISR costs (upper diagram), distinguished for a low-grade and high-grade deposit, and ore zone available for the leach solution at 50 days after wellfield start-up (lower figure) dependent on spacing between injection and extraction wells in the wellfield. ....... 48

Figure 22: Electron exchange per metal dissolved dependent on the mineralization. [Note: The chemical structure of covellite CuS is similar to sphalerite ZnS.] ................................................................................. 49

Figure 23: Proposed flow sheet for the BIOMOre process .................... 51

Figure 24: Mass balance from METSIM for water and acid washing. ... 56

Figure 25: Mass balance from METSIM for bio-leaching and bio-oxidation. ............................................................................................... 56

Figure 26: Mass balance from METSIM for solvent extraction. ............. 57

Figure 27: Mass balance from METSIM for copper electrowinning. ...... 57

Figure 28: Mass balance from METSIM for Effluent Treatment Plant (ETP). .................................................................................................... 58

Figure 29: SX- EW flow sheet. .............................................................. 59

Figure 30: IX – EW flowsheet. ............................................................... 60

Figure 31: Copper price over the last few years. ................................... 63

Figure 32: Long term copper prices. ..................................................... 64

Figure 33: The cost price relationship in 2009$. ................................... 66

Figure 34: Mines commissioned in 2006 and 2007. .............................. 67

Figure 35: Mines commissioned in 2008 and 2009. .............................. 67

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List of tables

Table 1: Mines and projects mined by ISR [from (Seredkin et al., 2016)] .............................................................................................................. 18

Table 2: Copper extraction hydrometallurgical technologies (pressure oxidatin/leaching) .................................................................................. 30

Table 3 Hydrometallurgical copper extraction technologies (atmospheric oxidation/leaching) ................................................................................ 30

Table 4: Copper raffinate composition .................................................. 33

Table 5: Copper PLS composition ......................................................... 33

Table 6 Comparative stages needed to run CLC and BIOMOre processes .............................................................................................. 35

Table 7 Preliminary OPEX of every CLC process stage ....................... 35

Table 8: KGHM deposit’s sandstone layer. ........................................... 53

Table 9: Sandstone layer composition in the Rudna mine. ................... 53

Table 10: Copper minerals of the sandstone layer in the Rudna mine. . 54

Table 11: Copper recovery rate from the blocks during ferric leaching. 55

Table 12: Recovery of individual mineral species during the ore leaching. .............................................................................................................. 55

Table 14: Capital costs of the purification process using solvent extraction. .............................................................................................. 59

Table 15: Operating costs of the purification process using solvent extraction. .............................................................................................. 60

Table 16: Capital costs of the purification process using ion exchange. 62

Table 17: Operating costs of the purification process using ion exchange. .............................................................................................. 62

Table 18: Copper products and their value from different small-scale production processes. ........................................................................... 63

Table 19: Scrap contribution as percentage of total use of copper. ...... 66

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List of abbreviations/acronyms

a Annum (year) AEC Anion Exchange Capacity (e.g. of clays, IX resins. BPT Best Practice/Practicable Technology CAPEX Capital Expenditure CEC Cation Exchange Capacity (e.g. of clays, IX resins. CLC Cobre Las Cruces, S.A. (Spain). Mining company. Partner in

BIOMOre. CNRS Centre National de la Recherche Scientifique (National

Center for Scientific Research), France, Partner in BIOMOre. d Day D Darcy (old unit of hydraulic conductivity/permeability) EC Electroconductivity, usually measured in mS/cm or µS/cm (S

– Siemens) Eh Redox potential (of a solution) with reference to the hydrogen

electrode. Cf. ORP EIA(EIS) Environmental Impact Assessment (Statement) EPS Extracellular Polymeric Substance ENA Enhanced Natural Attenuation ETW Effluent Treatment Plant EW Electrowinning FEFLOW (Software) Finite Element subsurface FLOW system.

Computer program for simulating groundwater flow, mass transfer and heat transfer in porous media and fractured media

FIGB Ferric Iron Generating Bioreactor FLT Field Leach Trial GEIS General Environmental Impact Statement GT Grade x Thickness [wt%ꞏm], cf. Productivity (Glossary) GW Groundwater h Hour HF Here: Hydraulic Fracturing. HRT Hydraulic Retention Time

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IMN (Instytut Metali Nieżelaznych) Institute of Non-Ferrous Metals, Gliwice, Poland

ISF Imperial Smelting Furnace ISL In-Situ Leach(ing), also referred to as In-situ Recovery ISR ISR In-Situ Recovery, also referred to as In-situ Leach(ing) ISL IX Ion eXchange kg Kilogramm KGHM (Former Kombinat Górniczo-Hutniczy Miedź). Polish Mining

Company (nowadays KGHM Polska Miedź SA) and partner in the BIOMOre project.

KiLea-Hy (Generic) Kinetic Leach Model Software for ISR feasibility studies and the prediction of ISR production rates in dependence on deposit parameters and wellfield design. Extended to block leach applications within the BIOMOre project, i.e. extended to various Hydrological scenarios.

km Kilometer kWh Kilowatt hour L (or l) Liter LME London Metal Exchange LSR (sometimes L/S or S/L) Liquid to solid ratio (e.g. leachant

mass per ore mass in leaching operations) m Meter M Mole (mM – miliimole) MARP Mining And Rehabilitation Program METSIM (Software) Steady-State & Dynamic Process Simulator for

chemical and metallurgical processing mg/L Millgrams per liter (also referred to as ppm – parts per

million) MNA Monitored Natural Attenuation NA Natural Attenuation NF Nanofiltration NPV Net Present Value (model for economic assessment) OECD Organisation for Economic Co-operation and Development OPEX Operational Expenditure

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ORP Oxidation-Reduction Potential (as measured with reference to dedicated electrodes). Cf. Eh.

PER Public Environmental Report PFD Process Flow Diagram pH pH value (negative decade logarithm of hydrogen ion

concentration in mol/L) PLC Programmable Logic Controller PLS Pregnant Leaching Solution PVE Pore Volume Exchange ppb Parts per billion ppm Parts per million RLE Roasting-Leaching & Electrowinning RO Reverse Osmosis ROM Run Of Mine (ore in its natural, unprocessed state for

processing) s Second SAGB Sulfuric Acid-Generating Bioreactor SDI Sustainable Development Indicator SX Solvent eXtraction t Ton/Tonne (1,000 kg) TDS Total Dossolved Solids WACC Weighted Average Cost of Capital WF Wellfield (system of injection/extraction wells) WP Work Package (of BIOMOre project) wt% Weight percent 2D and 3D Two-dimensional/three-dimensional

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List of annexes

Annex 1: Glossary of terms Annex 2: Process flow diagram for the main process steps of the

BIOMOre process; pretreatment by water/acid washing, in-situ (bio-) leaching, solvent extraction, electrowinning, and effluent treatment.

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Executive summary

The CAPEX/OPEX analysis of the BIOMOre ISR technology has been separated into the three fundamental stages:

ISR operation itself, i.e. including wellfield development, construction and oper-ation as well as leachant refortification (recovery of the metal of interest from the ore body, thus, resulting in a PLS for further processing).

The separation of the metal(s) of interest from the PLS, either by ion exchange (IX) or solvent extraction (SX).

Metallurgical refining to produce the pure metal (usually by electrowinning).

Whereas the BIOMOre ISR technology is innovative (subject to validation of viability), the metallurgical separation and processing are industrial standard (however, various options could be applied that meet state-of-the-art technology criteria).

Regarding the primary recovery (ISR), the effect of both deposit-dependent and oper-ational parameters on the specific cost figures (€ per kg metal) of an ISR process was quantified with reference to economic criteria. The production rates simulated by the generic model software KiLea-Hy (cf. deliverable D7.1) were combined with an eco-nomic model to calculate CAPEX/OPEX for a quantitative assessment of an acid ISR operation applied to reduced (sulfidic) ore. These combination allows the quantification of specific ISR costs (in €/kg metal) dependent on deposit-related and operational pa-rameters. The following feasibility criteria were systematized with reference to specific costs figures (including both CAPEX and OPEX per kg metal):

Abundance of calcareous minerals in the ore formation limiting the applicability of acidic ISR to levels < 2 wt% (critical already at 1 wt% in many cases).

Ore grade (lower limit dependent on metal price).

Depth of the ore deposit (limit also dependent on metal price – for Cu about 200 m under current market conditions, but could be up to 1,000 m for high value metals).

Porosity-permeability (in most cases limited to porosity > 10 % and permeability > 0.1 m/d or 120 mD). Fracturing could enhance permeability in certain limits (i.e. increasing the flow rate), but could not improve porosity (as a measure of the available reaction space of effective solid/liquid ratio) due to the incompress-ibility of rock.

Spacing (typical distance between injection and extraction wells) at an eco-nomic optimum (usually between 30 and 50 m) that is also influenced by the ore grade and the value of the metal.

Specific oxidation potential to leach the metal (thus, determining the costs to oxidise the leachant before injection). This cost factor is dependent on both the electron balance of redox leaching and the impact of competing reductants.

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The software tool METSIM has been applied to the complete BIOMOre processing (IX/SX followed by electrowinning) considering: (i) water & acid washing, (ii) leaching & bio-oxidation, (iii) Cu extraction from PLS, (iv) Cu electrowinning, and (v) effluent treatment. The process flow diagrams were the basis for the CAPEX/OPEX estimates of the BIOMOre process including an economic comparison between metal recovery by SX and IX.

ISR operation costs were combined with the subsequent hydrometallurgical pro-cessing. Therefore, CLC extended the process flow model in METSIM to simulate the complete BIOMOre process inclusive the new stages of water and acid washing ap-plied in the Rudna Mine block reactor. The stages were required to avoid (i) high levels of chloride feeding the bacterial culture, and (ii) precipitation of hydrous ferric oxides in the main leaching stage.

Using above mentioned information the Material Balance for the process has been developed producing enough information to estimate process reagents and utilities consumptions to be integrated in the operating expenditures (OPEX). Labour and maintenance cost have been estimated using average data from similar industrial ap-plications to finally calculate the complete BIOMORE OPEX for the process plant.

With the METSIM model as reference, the total investment cost has been developed by means of the conceptual engineering study of each process stage. CAPEX esti-mates have been developed using stochastic estimating methods such as cost/capac-ity curves and factors, scale of operations factors, Lang factors, Hand factors, Chilton factors, Peters-Timmerhaus factors, Guthrie factors, and other parametric and model-ling techniques. The comparison showed that SX is economically advantageous.

In summary, every deposit and every ISR application is specific. However, the main trends of feasibility criteria for the economic application of the BIOMOre technology have been deduced. Reliable software tools have been developed (KiLea-Hy) or suc-cessfully configured (METSIM) to simulate the technological parameter and complete cost figures for a BIOMOre ISR application.

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Content

List of figures ........................................................................................... 5

List of tables ............................................................................................ 7

List of abbreviations/acronyms ................................................................ 8

List of annexes ...................................................................................... 11

Executive summary ............................................................................... 12

1. Introduction .................................................................................. 16

1.1. CAPEX/OPEX analysis – the general framework ........................ 16

2. Review of state-of-the-art mining technologies ............................ 17

2.1. In-situ recovery technology .......................................................... 17

2.1.1. Overview of ISR applications ....................................................... 17

2.1.2. Actual ISR applications for copper ............................................... 18

2.1.3. Copper ISR operation and associated costs ................................ 21

2.2. In-situ recovery technology – an overview ................................... 23

2.2.1. State-of-the-art ISR application .................................................... 23

2.2.2. ISR performance and cost estimates ........................................... 25

2.3. State-of–the-art hydrometallurgical processing of copper ore ..... 27

2.3.1. Overview of hydrometallurgical copper processing applications .. 27

2.3.2. Hydrometallurgical copper processing in the CLC process ......... 32

2.3.3. Cu ores processing and associated costs .................................... 34

2.4. Study on zinc ore processing. ...................................................... 36

2.4.1. State-of-the-art technologies for Zinc recovery ............................ 36

2.4.2. Economic figures .......................................................................... 38

3. CAPEX/OPEX analysis ................................................................ 39

3.1. Cost estimates for copper ISR ..................................................... 39

3.1.1. ISR cost factors (CAPEX, OPEX) ................................................ 39

3.1.2. CAPEX/OPEX cost structure for ISR projects .............................. 40

3.1.3. Oxidative vs. non-oxidative ISR of Cu .......................................... 43

3.1.4. Dependence of costs on the depth of the deposit ........................ 43

3.1.5. Systematics of factors influencing costs and economic feasibility 44

3.2. Cost estimates for copper purification .......................................... 50

3.2.1. Cost effective flow sheet .............................................................. 50

3.2.2. Main streams and effluent compositions ...................................... 52

3.2.3. Mass balances ............................................................................. 54

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3.2.4. CAPEX and OPEX estimates Copper recovery from PLS ........... 59

4. Effects on copper prices ............................................................... 63

5. Cu ISR cost review and conclusions ............................................ 69

6. References ................................................................................... 72

7. Annex ........................................................................................... 74

7.1. Glossary of terms ......................................................................... 74

7.2. Process flow diagram for the main process steps of the BIOMOre process .................................................................................................. 78

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1. Introduction

1.1. CAPEX/OPEX analysis – the general framework

The CAPEX/OPEX analysis is a fundamental task within a (pre-)feasibility study for a mining project. Usually it is based on the detailed investigation of the deposit within the hydrogeological framework, the reliable estimate of resources (total figure) and reserves (technically and economically recoverable amount of metal) in accordance to international standards, the development of a viable technology including mining op-erations, metallurgical processing and all auxiliary/infrastructural conditions, finally re-sulting in a process flow diagram (PFD) defining all processing stages, the equipment, the mass flow balance and processing efficiencies, the consumables and residues.

In the case of the BIOMOre project, there are three fundamental stages:

ISR operation itself, i.e. including wellfield development, construction and oper-ation as well as leachant refortification (recovery of the metal of interest from the ore body, thus, resulting in a PLS for further processing).

The separation of the metal(s) of interest from the PLS, either by ion exchange (IX) or solvent extraction (SX).

Metallurgical refining to produce the pure metal (e.g. by electrowinning).

The ISR feasibility of the BIOMOre ISR process has been studied within WP7.1 in detail. The ISR model is suitable to simulate production scenarios for given deposit conditions, wellfield design and operational parameters, finally resulting in (time-de-pendent) recovery curves, detailed figures of chemicals’ consumption, and further op-erational data relevant for the cost analysis. This is the basis for the estimate of both CAPEX and OPEX. The ISR model is linked to the corresponding economic model to provide the required data for assessing the viability of the BIOMOre technology.

Both separation (IX or SX) and metal refining are considered to be state-of-the-art technologies, however, subject to select the most efficient and cost-effective techno-logical options.

When developing a new technology, it is essential to take into account the status of current technologies. The present study makes reference to the representative cost figures of European copper mining companies.

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2. Review of state-of-the-art mining technologies

2.1. In-situ recovery technology

2.1.1. Overview of ISR applications

ISR has been applied to many different commodities, namely uranium, copper, gold, nickel, scandium, rhenium, rare earth elements, yttrium, selenium, molybdenum, and vanadium. In 1959, ISR of uranium began and is now applied in many countries. It reached 51% of the world uranium production in 2014. For copper, ISR was introduced in the 1970s with several successful pilot tests and mines. Mining of scandium, rhe-nium, rare earth elements, yttrium, selenium, molybdenum, and vanadium as by-prod-ucts of uranium extraction was performed in pilot tests. Over the last years, ISR of gold, copper, nickel, rare earth elements, and scandium has been successfully devel-oped (Seredkin et al., 2016).

Further research was done on additional elements using sulfuric acid solutions or other solvents, with or without additional oxidants. An overview of ISR mines and pilot tests is given in Table 1.

The application of ISR for many different elements has led to some advances, includ-ing (Seredkin et al., 2016):

Improved technology:

o Use of a variety of leaching reagents, not only sulfuric acid and bicarbonate, but also thiosulfate, hypochlorite, and acidified chlorine, as well as thio-urea instead of cyanides for gold leaching, and sulfites of sodium or potash for other elements.

o Use of oxidants: oxygen gas, hydrogen peroxide, sodium peroxide, sodium nitrite, or nitric acid.

o Leaching elements below the water table by conventional filtration method and above the water table from unsaturated rocks using the infiltration method with collection of pregnant solutions at the surface of the water table.

o Use of natural permeability as well as the creation of artificial permeability by blasting and hydro-fracturing (e.g. Mount Isa).

Application to new types of deposits amenable to ISR.

Application to secondary (“technogenic”) resources (tailings, ashes, flooded un-derground mines, etc.).

Increasing dissemination of knowledge gained from the use of uranium ISR.

Despite the increase in uranium ISR applications, the use of ISR to recover other com-modities is still not common in the mining industry due to insufficient know-how and experience in ISR (a technology that is quite different from conventional mining

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operations) and a lack of understanding surrounding the geological environments amenable to ISR implementation (Seredkin et al., 2016).

Table 1: Mines and projects mined by ISR [from (Seredkin et al., 2016)]

2.1.2. Actual ISR applications for copper

The application of ISR for copper leaching has been mainly investigated and used for deposits in Arizona, USA. Most copper ISR applications are in porphyry deposits. These are globally abundant and the mineralization process results in a high density of uniformly distributed mineralized fractures; weathering may result in soluble copper mineralization (Sinclair and Thompson, 2015). Although ISR is rarely used for sulfidic copper ores, it is feasible by regenerating ferric iron using oxidants such as atmos-pheric bacteria cultures, pressurized oxygen gas, or chemical oxidants (Sinclair and Thompson, 2015). Copper is the most common base metal extracted by bioleaching (~15% of worlds copper production), mainly applying heap (bio-)leaching (CSIRO, 2015), but until now, Cu ISR has not been combined with bioleaching. A compilation of the main copper ISR applications is given in the following:

The Gunnison copper ISR project in Arizona, owned by Excelsior Mining Company, is the most promising project for copper ISR. It is located in southern Arizona, 65 miles

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east of Tucson, in a desert environment and relatively far away from populated areas. It would be the first greenfield project to apply an ISR method for copper in the USA. An updated prefeasibility study was published in February 2016. The host rock of this project (oxidized copper deposit – porphyry deposit) is solid and impermeable, but also naturally highly fractured. The majority of the copper mineralization is located on the surface of the open fracture veins. Extraction via a closed loop system of delivery and recovery wells is proposed and the pregnant solution will be processed in a Solvent-Extraction Electrowinning plant. Beginning in 2018 and spanning a 20-year mine life (or even longer - some mining information is given for 27 years), Gunnison Copper is projected to produce more than 1.6 billion pounds of copper and will add more than US$2.94 billion to the Gross State Product of Arizona, while operating with little to no noise, dust, or emissions.

The Florence Copper project in Arizona is a well-developed project. It was acquired by Taseko in 2015. In contrast to the Gunnison copper ISR project, it is located close to civilization and, thus local opposition is attempting to block the permitting process. At present, two outstanding permits are required to move forward with the Phase 1 Test Facility. This project has been under investigation from 1992 until present and commercial operation is proposed. It is a shallowly buried porphyry copper deposit in a groundwater saturated formation. It has probable reserves of 340 million tons grad-ing 0.358 % Cu containing 2.42 billion pounds of copper. The project obtained all major environmental permits under BHP in the late 1990s; currently, updating is being per-formed. A prefeasibility study from 2013 announced a mine life of 25 years and test work results showed that ~70 % copper recovery is possible (Sinclair and Thompson, 2015).

The Van Dyke project (Copper Fox Metals) is a 1,100 acre oxide copper deposit (porphyry deposit) located approximately 1.2 miles south of the Van Dyke oxide copper deposit in the Globe Miami mining district, app. 90 miles east of Phoenix, Arizona. It has been investigated from 1976 until present. Preliminary economic assessments were already performed (2015) and the calculated total copper grade was averaged at 0.76%. Further, in-situ pressure leach tests showed that 89% of the copper is solu-ble (acid and cyanide soluble).

The San Manuel copper mine (BHP Billiton) is located in south-eastern Arizona. Here, open pit and underground mining were combined with ISR using weak sulfuric acid to leach the oxidized portion of the porphyry mineral resource. The PLS was re-covered either in collection areas in the underground workings or in nearby recovery wells (Carlson et al., 2000). The oxide cap at the surface of San Manuel was, to a large extent, removed by open pit mining. ISR was applied to the exposed mixed oxide and sulfide mineralization zone that transitions to primary sulfide ore with increasing depth. ISR was used to leach remnant material in the pit walls, around underground workings, and on low-grade dumps (Seredkin et al., 2016). The deposit is unsaturated with groundwater (Sinclair and Thompson, 2015). The San Manual copper mine produced approximately 3.25 billion pounds of copper during 14 years of production.

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The Santa Cruz Pilot Plant, in Pinal County, Arizona (until 1998, Joint Venture be-tween ASARCO & Freeport McMoRan) produced Cu cathodes with a purity of 99.99% and a solvent extraction feed containing 0.3 to 1.8 g/L Cu by conventional mining. Nowadays, it is owned by Amrich Minerals, LLC, who plan to extract copper by ISR. The Santa Cruz Deposit is a porphyry copper deposit in a groundwater saturated for-mation (ore zone: 2,000 m long, 800 m wide, 1,600 m deep, thickness 1,700 m) (Amrich Minerals, 2013; Sinclair and Thompson, 2015).

Furthermore, ISR was applied for Bisbee (Briggs and Allison, 2015), the Brownfield copper project, Casa Grande, Mineral Park, Kimbley, Nacimiento, and Mina Sur.

There are some examples where the application of ISR of prior fractured rock was studied. For example, Rosenbaum & McKinney, 1970, studied a concept for in-place leaching of a nuclear fractured, deeply buried copper sulfide deposit (Safford deposit, Arizona) using an acid-ferric sulfate solution. At the Miami Copper Mine in Gila County, Arizona, the leach solution percolates through 600 ft of broken rock to a collection reservoir in the old mine (in-place solution mining). Afterwards, the PLS is pumped to a surface plant for cementation of the copper on shredded iron cans (Rosenbaum and McKinney, 1970). The in-situ leaching project at Miami was run for 7 years and closed down in 1982 (Seredkin et al., 2016).

Other projects involving in-place solution mining of broken and fragmented rocks, e.g. by blasting were: Ray (1937-1961), Tyrone (1941-1949), Pilares (1946-1960), Inspira-tion (1965-1974), Lakeshore (1983-1994), San Manuel (1995-2002), the Old Reliable mine (1972-1981) near Mammoth, Arizona, the Zonia project (1973-1975) located south of Prescott, Arizona, and the Big Mike mine (1973-1979) in Pershing County, Nevada (Briggs and Allison, 2015). Further, ISR trials were carried out at several sites near Mount Isa, Queensland in the 1970s. At the Gunpowder Mine, ISR was applied to broken ground (fragmented through blasting) beginning in the early 1990s (Seredkin et al., 2016).

An example of an ISR project outside of Arizona is the Gumeshevskoye Copper De-posit in the Ural Mountains of Russia. It was mined over nearly 300 years and it was proposed to recommence open pit mining of the oxidised mineralisation. However, feasibility studies showed this to be unprofitable. Instead, an ISR mine was selected and has operated since 2004 with an annual production of 5,000 t of copper. The cop-per grade in the pregnant solutions is 1-4 g/L with recovery of 49-120 % (recovery >100 % due to leaching copper outside estimated mineralised bodies) (Seredkin et al., 2016).

In summary, ISR has primarily been applied for the supplemental recovery of copper from established open pit and underground mines using sulfuric acid or acidic ferric sulfate solutions. For copper ISR, the major economic challenge is the limited expo-sure of copper minerals to the leach solution in complex subsurface flow paths, which has generally resulted in lower recoveries compared to established processing meth-ods (Sinclair and Thompson, 2015). Copper ISR operations have reported estimated recoveries in the range of 20 – 70 %, whereas conventional heap leaching generally

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results in 65 – 90 % recovery. Copper grades up to 4 g/L are reported in the pregnant leach solution from oxidized copper deposits.

2.1.3. Copper ISR operation and associated costs

The technical requirements for in-situ leaching in the BIOMOre process are described in detail in the BIOMOre deliverable D7.1. In the following, the specific costs are dif-ferentiated by cost types. Costs of the initial phase of a mining project, e.g. land ac-quisition, exploration, technical studies, initial environmental permitting, as well as gen-eral expenses, e.g. marketing and administration, are comparable for all mining pro-jects and not considered here.

Relevant CAPEX of an ISR operation are (Hackman, 1982; Pugliese, 1989):

Well site preparation costs for injection and production wells, stimulation wells, and monitoring wells:

Drilling, well casing, cementing, completion, and stimulation costs.

Well logging costs for exploration, especially for horizontal drilling and stim-ulation.

Stimulation equipment, if required.

Well injection and production equipment costs including well houses, e.g. pumps, automation/control units.

Leach solution transfer system:

Equipment for the refortification of leach solution (including FIGB in the BI-OMOre approach).

Monitoring/dosing/drainage systems.

Infrastructure e.g. road construction, power supply, pipe lines.

Wellfield restoration costs.

Relevant OPEX of an ISR operation are (Hackman, 1982; Pugliese, 1989):

Personnel, supervision and technical support.

Maintenance and repair.

Chemical and consumable costs (e.g. sulfuric acid, oxidant).

Infrastructure, e.g. electric power, water, fuel.

Transportation.

Security.

Cost of utilities.

The Cu separation costs (IX/SX) are considered separately.

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The main target for the wellfield operation is a constant, preferably high, production rate. Strong fluctuations in the production rate of the wellfield operation increase the costs of the subsequent process steps. Thus, the wellfield operation is a combination of the technical operation and the wellfield production management. Usually, the well-fields are developed and commissioned in a subsequent manner in order to realize an approximately constant production rate (i.e. lower production from wellfields already depleting are compensated by the consecutive start-up of new wellfields (sometimes individual extraction patterns).

In addition, to the costs required to operate the wellfield for production, wellfield related remediation costs have to be taken into account. These depend on requirements from a regulatory authority, with three different options for wellfield remediation available:

Natural attenuation (NA), if underground conditions are suitable to compensate the reactivity of the leaching solution by the long-term interaction with the host rock formation.

Enhanced natural attenuation (ENA), i.e. neutralizing or reducing reagents are added to the mining aquifer to expedite NA. Biochemical/biotechnological meth-ods have also been developed and applied successfully (subject to nutrient ad-dition to increase the remediation effect catalysed by microorganisms).

Active treatment by pump and treat (required in sensitive hydrological systems) to re-establish baseline conditions.

In the case acidic ISR of metals from reduced minerals, main attenuation mechanisms are rock-water interactions including:

Neutralization due to (fast) dissolution of calcareous minerals, (long-term) dis-solution of clay minerals, feldspar, or others, finally causing the immobilization of metals that were leached during the active ISR phase.

Chemical reduction by sulfide minerals or organics accelerated by microbial ef-fects, e.g. by sulfate reducing bacteria. Lowering the redox potential creates another specific condition for immobilizing leached metals.

Sorption processes, e.g. ion exchange or surface complexation on clay miner-als, organic matter, etc.

The required measures for remediation may range from installation of monitoring wells to extending the wellfield operation by a period with low production rate to installing a water treatment plant. Thus, the amount of costs increases from NA to active treat-ment; however, the time of remediation decreases (cf. Figure 1).

D7.2 |

Figure 1: Remediation options and their effect on the post-mining phase

2.2. In-situ recovery technology – an overview

2.2.1. State-of-the-art ISR application

The technological scheme of ISR is illustrated in Figure 2 (explained in more detail in deliverable D7.1). It demonstrates:

The typical setup of wellfields consisting of injection and extraction wells, both screened in the mining zone to realize the leachant flow in a controlled regime.

The usual principle of consecutive development and operation of wellfields to stabilize the production rate at an average target level.

The main areas/points where energy (electrical), chemicals and water are con-sumed for pumping, leachant refortification/conditioning and processing.

The options of bioleaching considered in the BIOMOre project:

o Ex-situ biooxidation of Fe2+ to Fe3+ (ferric acting as direct oxidant to selec-tively leach the metal from the reduced mineral in the ore),

o In-situ bioleaching (option of O2 injection to stimulate the microbially cata-lysed oxidation of Fe2+ to Fe3+). In this case, the condition (constraint) of the maximum concentration of O2 as function of pressure and temperature applies).

Overall leachant cycle with minor bleed (0.5-2 % usually) to better balance well-field hydrology (not shown in this Figure).

D7.2 |

Figure 2: Schematic of ISR, in particular, emphasizing the consumption of chemicals, energy

and water (C, E, W, respectively). The BIOMOre options (in red circles) include ex-situ biooxidation and the potential in-situ bioleaching (both requiring O2).

D7.2 |

The main principles of ISR and the dependences of production rates (flow rate times metal concentration in the PLS) on both deposit parameters and operational parame-ters (wellfield design and operation) have been discussed in deliverable D7.1 and ap-plied to Cu ISR by using the generic software KiLea-Hy.

2.2.2. ISR performance and cost estimates

The ISR feasibility criteria have been categorized in the complexes ore morphology (including grade distribution), hydrogeology, mineralogy, groundwater chemistry, and microbiology in D7.1. Figure 3 demonstrates how these feasibility criteria are linked to the wellfield design and setup of the (bio-)leaching chemistry to realize a pore volume exchange (PVE) rate in correspondence to the leaching rate, both determining the production rate as function of wellfield operation time.

In addition, Figure 3 shows the link of the technological scheme to the economic model. The economic model is applied in two typical stages:

First the estimation of the CAPEX and OPEX figures for a given ISR production scenario. The comparison of the specific total costs (€/kg) versus the market price of the metal is already quite conclusive. KiLea-Hy has been extended to estimate CAPEX/OPEX figures in dependence on the deposit and operational conditions. The results from a systematic cost analysis for ISR operations are presented in Section 3.1.

Second the overall ISR mining project can be assessed by applying the net present value (NPV) approach to demonstrate, if the project is likely to be viable. With regard to the NPV model for ISR there are the following specific aspects that need to be taken into account:

o Due to the consecutive development and operation of wellfields throughout the mining project a considerable part of the CAPEX (i.e. for wellfield devel-opment and construction) is distributed over the lifetime of the mining pro-ject. The advantage is demonstrated in Deliverable D7.3.

o ISR planning and performance cannot be predicted in every detail, thus, re-sulting in ISR specific risks. ISR is an “invisible” mining technology, i.e. it is dependent on the high quality of exploration and delineation data as well as other relevant studies (hydrogeological/mineralogical) on the one side and the level of relevant 3D models to represent the data and to predict the ISR wellfield operation reliably. Significant progress has been made in the past years to develop the tools for establishing the ISR 4.0 vision (cf. D7.1).

D7.2 |

Figure 3: ISR feasibility criteria, wellfield design and performance and link to the economic

model (D7.1)

D7.2 |

2.3. State-of–the-art hydrometallurgical processing of copper ore

2.3.1. Overview of hydrometallurgical copper processing applications

The historic development of copper hydrometallurgy in the modern era has achieved the following main milestones:

Oxides/roasted cinders heap leaching. Early 1900´s.

Iron cementation applications. Early 1900´s.

Direct electrowinning. Early 1900´s.

Roasting, leaching, and electrowinning integration. 1960-70´s.

Leaching, solvent extraction, and electrowinning (L/SX/EW). 1970-80´s.

Enhanced sulfide heap (bio)leaching. 1990-2000´s.

Concentrate pressure leaching. 2000´s.

Sulfide ores and concentrates atmospheric leaching. 2000-2015.

Figure 4 summarises the up-to-date hydrometallurgical copper extraction technologies dependent on the ore type.

Figure 4: Copper extraction through hydrometallurgical processing

Copper oxide ores ([1] in Figure 4). The copper bearing oxidised ores are easily leached in acidic media; therefore, they are processed in dump leaching or heap leach-ing spread areas irrigated with diluted sulfuric acid or by using ISR, as performed in Arizona, USA. During heap leaching, the copper metal is released into solution,

D7.2 |

percolates through the pad, and is finally collected in dedicated ponds. In an ISR ap-plication, the released copper is pumped to the surface. The copper pregnant solutions are treated by means of solvent extraction and electrowinning (Cu SX-EW), obtaining copper cathodes. Nowadays, this approach is used broadly in copper mines in Chile and the USA. In ISR applications, the metal recovery from the pregnant solution by solvent extraction (SX) may be replaced by ion exchange (IX) to avoid organic medium in the leach cycle.

Copper sulfide ores ([2], [3] in Figure 4). The sulfidic ores require chemical or biolog-ical oxidation to leach and facilitate copper metal dissolution. For secondary copper sulfides, e.g. chalcocite or covellite, heap leaching or bioleaching is usually applied for low-grade copper ores, while atmospheric leaching in stirred tanks is utilised for high-grade copper ores, as in the Las Cruces mine (Spain) and Sepon mine (Laos). For chalcopyrite ores, which are more difficult to leach, new bioleaching technologies are currently under development in some places, such as the Escondida mine (Chile). Furthermore, ISR is being studied for the leaching of sulfide ores, as is the case in the BIOMOre project. Once the copper metal goes into solution, solvent extraction and electrowinning techniques (Cu SX-EW) are employed to yield copper cathodes.

Copper concentrates ([3] in Figure 4). Typically, commercial copper concentrates contain mainly chalcopyrite and bornite minerals and are treated through copper smelt-ing (e.g. in flash furnace) and further electro-refining techniques. At the end of the process, pure copper is collected on cathodes. It is remarkable that over 70 % of the worldwide copper metal production is obtained through smelting and electrolytic refin-ing processing. Large and modern copper smelters include Codelco Norte (Chile), Jiangxi Copper Corp. Guixi (China), and Aurubis Hamburg (Germany), all of them us-ing an Outokumpu flash furnace.

Nowadays, over 25% of the world’s copper production is derived from hydrometallurgy using sulfate, sulfate/chloride, or chloride media combined with SX-EW processes. In general, hydro-technologies allow over 95% metal recovery from selected copper ores or concentrates. Typically, oxidation is carried out through ferric ions, and the Fe3+ concentration is maintained, e.g. by oxygen injection, running at atmospheric or pres-sure leaching conditions.

Several technologies based on sulfate or mixed sulfate/chloride solutions and using stirred tank leaching conditions are currently operating at industrial scale or at different development stages, as shown in

D7.2 |

Table 2 and Table 3.

D7.2 |

Table 2: Copper extraction hydrometallurgical technologies (pressure oxidatin/leaching)

PRESSURE OXIDATION / LEACHING

Process Status Temperature Notes

High Pressure Commercial 200-230 ºC Fine grinding

FCX Commercial 150 ºC Fine grinding

Mount Gordon Commercial 90 ºC Plant shut down

CESL Demo plant 140-150 ºC Chloride, surfactant

Platsol Pilot plant 220-230 ºC Chloride

Dynatec Pilot plant 150 ºC Coal

AA-UBC Pilot plant 150 ºC Fine grinding, surfactant

Activox Pilot plant 110 ºC Fine grinding

Table 3 Hydrometallurgical copper extraction technologies (atmospheric oxidation/leaching)

ATMOSPHERIC OXIDATION / LEACHING


Sepon Copper Commercial 80 ºC Direct ore leaching

Cobre las Cruces Commercial 90 ºC Direct ore leaching

Albion Pilot plant 80 ºC Fine grinding

Galvanox Pilot plant 80-90 ºC Pyrite catalysed

Sical Development 80-90 ºC Silver catalysed

BIOOXIDATION / BIOLEACHING


BioCop Demo plant 70-80 ºC Plant shut down

Bactech Demo plant 40-50 ºC Fine grinding

D7.2 |

During copper leaching, other metals are extracted from the ore that may be a viable by-product. An overview of a technology for the recovery of copper and other metals is also given in Figure 5 and explained in Deliverable D4.2.

Figure 5: Scheme of proposed technology for copper and other metals recovery by hydro-

metallurgy methods.

Membrane separation techniques, such as nanofiltration (NF) and reverse osmosis (RO), can also be used for the pre-concentration of the PLS to increase the effective-ness of the extraction processes. As NF is the most suitable amongst pressure-driven membrane technologies due to process flexibility and cost advantages over RO, the focus is set to NF. An issue for the applicability of NF technique is scaling and fouling. These can be classified into:

Membrane scaling.

Metal oxide fouling.

Membrane plugging.

Colloidal fouling.

Biological fouling.

D7.2 |

For example, in the BIOMOre pilot test case, gypsum scaling would be the major lim-iting factor for NF. A pre-treatment of the solution that feeds to NF, as well as cleaning of the membrane, can help to avoid scaling and fouling (see Deliverable D4.4).

Many processes have been tested over the last years, but only a few have been com-mercialized, with variable success, due to difficulties achieving technical-economic vi-ability.

Only two commercial atmospheric leaching plants operate worldwide:

Sepon, MMG, Laos.

Cobre Las Cruces (CLC), First Quantum, Spain.

2.3.2. Hydrometallurgical copper processing in the CLC process

The CLC process is briefly described as a reference for the assessment of the viability of BIOMOre.

The CLC mine and integrated hydrometallurgical plant is located approximately 20 km NW of Seville in southern Spain. The plant has a nominal capacity of 72,000 t/y copper cathodes, grade A LME. High grade copper ores (>5% Cu) containing mainly chalco-cite mineralization are processed through crushing, milling, atmospheric leaching, sol-vent extraction, and electrowinning (cf. Figure 6). The obtained leaching residue is filtered, washed, and disposed of under dry conditions.

Figure 6: CLC process simplified scheme

Crushing & Milling: The aim of this stage is to increase the surface of the ore to enable chemical leaching in stirred tanks. Comminution comprises a dry (crushing) and a wet (milling) operation to comminute the run of mine (ROM) material with a

D7.2 |

characteristic size of 100% passing 1,000 mm to a ground material with a characteris-tic P80 over 100 microns. The crushing plant consists of a primary jaw crusher, a sec-ondary cone, and a tertiary cone in closed circuit. The milling plant includes a ball mill running in a closed circuit with corresponding hydrocyclones.

Atmospheric Leaching: This stage is the key of the CLC process where copper is leached in ferric sulfate media with oxygen sparging. Ferric sulfate is conveyed within the recycled copper raffinate from the SX circuit and mixed with ground ore in stirred tanks. A typical composition of copper raffinate with regards to the main components is shown in Table 4.

Table 4: Copper raffinate composition

Component Concentration (g/l)

Cu 10-14

Fe Total 45-50

Fe(II) 22-25

Fe(III) 22-25

H2SO4 55-65

The leaching process is carried out at atmospheric pressure and at temperatures around 90ºC. The following main reactions are involved:

CaCO3 + H2SO4 → CaSO4 + 2 H2O + CO2

Cu2S + 2 Fe2(SO4)3 → 2 CuSO4 + S0 + 4 FeSO4 (Chalcocite)

CuS + Fe2(SO4)3 → CuSO4 + S0 + 4 FeSO4 (Covellite)

S0 + 3 Fe2(SO4)3 + 4 H2O → 6 FeSO4 + 4 H2SO4

2 FeSO4 + H2SO4 + ½ O2 → Fe2(SO4)3 + H2O

FeS2 + H2O + 3 1/2 O2 →FeSO4 + H2SO4

The PLS is separated from the leached slurry in a thickener, cleaned by filtering and washing, and finally fed to the Cu-SX circuit to recover the copper. The typical compo-sition of copper PLS with regards to the main components is shown in Table 5.

Additionally, the PLS contains dissolved calcium from the gangue and is saturated with gypsum. Calcium is removed by gypsum precipitation. Therefore, the PLS passes through cooling towers, dropping the temperature to 35ºC. It then flows from the tower’s outlet to a thickener, where gypsum is removed in the underflow.

Table 5: Copper PLS composition

D7.2 |

Component Concentration (g/l)

Cu 35-40

Fe Total 40-45

Fe(II) 20-25

Fe(III) 20-25

H2SO4 20-25

Solvent extraction: The PLS from the leaching stage contains primarily copper and iron. The aim of the SX stage is to selectively extract copper from the PLS. Therefore, the two immiscible liquid phases are mixed in a counter-current regime, (i) an aqueous phase (PLS), and (ii) an organic phase bearing a chemical extractant (aldoxime in kerosene as diluent) that bonds copper. There are two counter current extraction cir-cuits, (i) a primary circuit for the primary PLS coming from the PLS pond, and (ii) a secondary circuit for the secondary PLS mainly composed of the belt filters washing outlet stream. Only one organic loop exists that is in contact with both circuits. The copper is stripped from the organic phase into the feed solution (electrolyte) for the electrowinning process.

Electrowinning: The EW plant consists of 144 cells with 84 cathodes each. The cath-odes are made out of stainless steel and anodes of alloyed lead. Two different types of cells exist in the EW operation: scavenger cells and commercial cells. The electro-lyte feeding these cells (commercial electrolyte), is a mixture of lean electrolytes from both the scavenger and the commercial cells. The main reactions that occur in the electrolytic cells are:

Cathode: CuSO4 + 2 e- → Cu + SO42-

Anode: H2O → ½ O2 + 2 H+ + 2e-

2.3.3. Cu ores processing and associated costs

Copper ore beneficiation through bio- or hydrometallurgical methods usually requires a combination of process stages. Figure 7 compares the stages of conventional mining with an ISR process for uranium. In line with the prior scheme, a comparison of the equivalent process stages needed by the CLC and BIOMOre processes is presented in Table 6.

To estimate the operating and capital costs of a specific copper extraction process, it seems convenient to know the costs associated with every process stage. In this re-gard, Table 7 presents some preliminary costs of the CLC process that may be useful for BIOMOre project cost estimates.

D7.2 |

Figure 7: Copper extraction simplified scheme compared to the U ISR process.

Table 6 Comparative stages needed to run CLC and BIOMOre processes

CLC Process BIOMOre Process

Exploitation. Ore Mining Drilling

Crushing & Milling Hydro fracturing & Stimulation

Atmospheric Stirred Tank Leaching

- Oxygen sparging

In-situ Leaching and Pumping

- Fe(II) Bio-oxidation

Solvent Extraction-Electrowinning Solvent Extraction-Electrowinning

Ancillaries Ancillaries

Table 7 Preliminary OPEX of every CLC process stage

CLC Process stage Euro/t Cu

Comminution 150-210

Atmospheric Leaching 160-200

Solvent Extraction 50-90

Electrowinning 190-220

Effluent Treatment 5-10

Total 555-730

D7.2 |

Furthermore, the results of Deliverable D4.1 of WP4 will be taken into account where the CAPEX and OPEX of different recovery options were investigated. CLC has de-veloped a conceptual process flow sheet of the BIOMOre process and its mass bal-ance within WP4 that may be a suitable reference case for Task 7.2.

2.4. Study on zinc ore processing.

A literature review of zinc recovery and processing has been done, aiming to deter-mine the possible application of the BIOMOre concept to one of the alternatives to copper (cf. deliverable D7.1).

Currently, zinc processes focus on the production of a concentrate followed by RLE technology (Roasting-Leaching & Electrowinning). If lead is present in the concen-trates, the smelting process is used (cf. Figure 8Figure 8).

Figure 8: Zinc/Lead recovery routes.

The use of the BIOMOre concept needs to be focused on complex and low grade ores. A typical ore deposit to be used for techno-economic calculations of the proposed technology can then be defined.

2.4.1. State-of-the-art technologies for Zinc recovery

The majority of the world’s zinc production is performed using RLE technology. The feed material (zinc concentrate) comes from the mineral processing of zinc sulfide ores by means of froth flotation. Figure 9 shows a conceptual process scheme with the details involved in this technology.

By combining RLE and direct leaching of zinc concentrates, zinc can be efficiently produced from a sulfide concentrate (cf. Figure 10). This process has the advantage

D7.2 |

of sulfuric acid production from sulfur oxide and its consumption in the direct leaching process.

Figure 9: Direct leach process with combined ferrite/sphalerite leaching in one stage.

Figure 10: Combination of Zn current technology and direct leaching

D7.2 |

2.4.2. Economic figures

Depending on the zinc production technology, the associated operation and construc-tion costs may vary. In the case of zinc production, about 80 % of the zinc is produced using the electrolytic process with an operation cost of approximately 20 c/lb (euro/ton). Imperial smelting furnace (ISF) and electrothermic processes have an op-eration cost around twice as much as the electrolytic process. The different OPEX are presented in Figure 11.

Figure 11: Operating costs for zinc production dependent on the process time evolution.

For CAPEX, a new electrolytic smelter ranges from approximately US$1000 to 4000/t zinc capacity (in 2009$ terms) for a smelter with a minimum capacity of 100kt/a.

D7.2 |

3. CAPEX/OPEX analysis

3.1. Cost estimates for copper ISR

3.1.1. ISR cost factors (CAPEX, OPEX)

The feasibility criteria for ISR applications were discussed in detail in deliverable D7.1. In particular, the dependence of the leach performance (production rate) on various variables including both deposit characteristics and performance parameters was studied. This Section describes the economic assessment of ISR, again in depend-ence on deposit characteristics and ISR performance parameters (wellfield design, operation). Therefore, the generic Kinetic Leach software – KiLea-Hy (deliverable D7.1) has been linked to an economic model for ISR.

Every ISR project is specific. In addition to specifics in the regional hydrogeological system (overall stratigraphy, confinement conditions, etc.) and the deposit character-istics (total resource, ore grade, specific mineralogy, etc.) the costs (both CAPEX and OPEX) are considerably influenced by the socio-economic, regulatory, climatic and other factors.

In order to provide quantitative figures and dependences (trends), average conditions and representative cost figures in OECD countries have been assumed. Note that quite considerable deviations might apply, subject to adjustments for each individual ISR project in each individual country.

The following individual CAPEX items have been included:

Delineation (drillings and logging/assays to better characterize the ore body for an improved wellfield design). Dependent on depth of deposit.

Injection/extraction well construction (including drilling). Dependent on depth of deposit.

Monitoring wells. Dependent on depth of deposit.

Wellhouse construction. Reusable. 30 % amortization rate assumed.

Filter skid. Reusable. 30 % amortization rate assumed.

Chemical dosage station. Proportionate for wellfields.

Electrical/PLC. Reusable. 30 % amortization rate assumed.

Trunkline/pipework costs. Proportionate for wellfields.

OPEX items considered in the present study include:

General OPEX costs during wellfield lifetime (personnel, repair and mainte-nance, EHS – environment, health, and safety measures, etc.).

D7.2 |

Acid consumption (differentiated with regard to acidification of the relevant groundwater body, compensation of calcareous minerals, and refortification during the wellfield lifetime to compensate long-term neutralization effects).

Addition of a ferric compound like Fe2(SO4)3 to built-up Fe concentration to tar-get (dissolution of Fe minerals not sufficient in most application cases).

Oxidation of lixiviant (re-oxidation of ferrous to ferric Fe for quantitative oxidative leaching).

Energy consumption for pumping (depth dependent for extraction pumps).

Remediation costs (strongly dependent on regulatory requirements; here it is assumed that (E)NA, i.e. (enhanced) natural attenuation could be applied (in addition to well plumbing).

In order to calculate the cost figures for given ISR conditions (deposit, performance) on the basis of the production rate directly, the generic ISR software tool KiLea-Hy (cf. deliverable D7.1) has been linked to the economic model to calculate the total and the specific cost figures (€ per kg metal). The embedded database for basic cost figures corresponds to representative values derived from uranium ISR projects and feasibility studies. As already pointed out, these underlying cost data are subject to adjustments to individual ISR project conditions. However, this limitation does not influence the main results/trends derived and conclusions drawn.

3.1.2. CAPEX/OPEX cost structure for ISR projects

In order to demonstrate the typical cost categories for ISR projects, an copper ISR production scenario has been assumed that is summarized in Figure 12 (similar to the simulation options described in deliverable D7.1).

The main assumptions include:

ISR of a reduced Cu deposit at 250 m depth, further characterized by 25 % porosity, 1.1 m/d permeability, 3 wt% Cu mineralized as chalcocite (i.e. amount of electrons exchanged for the oxidative leaching is 5 mol e-/mol Cu), 0.5 wt% calcite.

7-spot wellfield at 11,700 m3 ore volume (5 m thickness) per wellfield pattern, containing 698 t of Cu.

Injected lixiviant at pH = 1.5 (H2SO4), containing 5.6 g/L Fe3+ (100 mmol/L elec-tron acceptor equivalent). No competing reductants in the ore body assumed at this stage (i.e. simulation of quite ideal conditions).

Operating at 28.8 m3/h per pattern (8 L/s), resulting in a pore volume exchange (PVE) rate of 0.261 d-1 or equivalent transient time of about 4 days.

Acid breakthrough achieved after 49 days, followed by the oxidative leaching of Cu.

Lixiviant cut-off concentrations assumed to be 100 ppm.

D7.2 |

Total wellfield operation during 1083 years (roughly 3.0 years) resulting in a recovery of 529 t Cu (76 % recovery).

The leaching is thermodynamically constrained (i.e. the leached Cu corre-sponds to the concentration of oxidant in the injected leachant in order to meet charge conservation in the redox reaction).

Figure 12: KiLea-Hy leach performance model – metal concentration in PLS and in-situ re-covery (left) and daily production and ore resource over the time (right) for the ISR scenario

described in the text.

The assumed production scenario results in the following total cost figures:

CAPEX: 0.48 €/kg (mainly caused by well construction: 0.38 €/kg).

OPEX: 2.81 €/kg (mainly caused by the oxidation of ferrous to ferric at the amount of 1.95 €/kg). Note that the equivalent costs for H2O2 as oxidant has been used, since reliable data for the ex-situ bio-oxidation of the lixiviant haven’t yet been available.

Total costs for ISR only: 3.29 €/kg.

Approximate (estimated) costs for metallurgical processing (SX or IX followed by electro-winning); moderate-scale plant: 1.46 €/kg (0.80 €/kg CAPEX, 0.66 €/kg OPEX).

Leach Kinetics and Hydrology - Generic KiLea_Hy Leach system: WF on/off 1 Block off/on 0

Parameter Input Comment Parameter Input CommentpH_inj 1.5 pH<2 acid / pH>7 alkaline Ore volume [m3] 11,700 Estimate from wellfiedl setup[Ae-] [mmol/L] 100 Ferric equivalent 5,600 mg/L Effective porosity 25.0% Free-fluid porosity

O equivalent 800 mg/L Dry ore density [t/m3] 1.99 Estimate from matrix density 2.65

[C] [mmol/L] 0 HCO3 equivalent 0 Ore grade [wt%] 3.00% Effective metal grade in model space

A_metal 64 Metal mass number Mass_metal [kg] 697,613 Total metal mass in model space [Ae-]/[metal] 5 mmol/mmol - redox factor Recovery factor 95% Effective fraction of leachable metalCalcite [wt%] 0.5000% m_recoverable [kg] 662,732 Leachable metal mass in model space

Access. calcite [wt%] 0.4564% ~(/0)d; 0=0.3 m_ISR-amenable (kg) 547,303 ISR amenable metal (-S-corrected)

Acid breakthrough 49 Days (effective time) V_p_tot [m3] 2,646 Effective pore volume in model spacer[1/d] 0.00374 Leachability Permeability [m/d] 1.1111 Estimate Kozeny-Carman-Equation

Maximum Q [m3/h] 53.049 Thiem equation (WF) or Darcy law (Block)Lix cutoff [ppm] 100 Condition for economic recovery Setup Q [m3/h] 28.800 From hydrological modeling/real dataAverage grade [ppm] 706 Q actual [m3/h] 28.800 Conditions: Q<Qmax and q<1WF lifetime 1083 days 10 years maximum 2.97 years q [1/d] 0.261 PVE rate: q = Q/V_p [per day]Recovery 529124 kg Av. transient time [d] 4

ppm

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D7.2 |

Total specific costs (ISR and processing): 4.75 €/kg.

The individual CAPEX and OPEX figures for the given production scenario are shown in Figure 13 and Figure 14, in particular, highlighting the dominant cost components. Note that this simulation was performed for quite ideal conditions (no redox competi-tors). Real oxidant expenditures could be double (i.e. 4.76 €/kg OPEX resulting in a total of 6.70 €/kg (ISR and processing).

Figure 13: CAPEX cost factors for oxidative Cu ISR production scenario described in the

text. Total specific CAPEX: 0.48 €/kg.

Figure 14: OPEX cost factors for oxidative Cu ISR production scenario described in the text.

Total specific OPEX: 2.81 €/kg.

The costs for the ex-situ biooxidation would be different compared to the dosage of H2O2 to simply oxidize the lixiviant directly into the trunkline supported by a static mixer (industrial practice). In this case, one has to consider the proportionate CAPEX for the FIGB (ferric iron generating bioreactor) plus applicable OPEX (pumps, blowers for aer-ation, nutrients, repair and maintenance). The costs depend on the hydraulic retention time (HRT) of the bioreactor. HRT in the order of 10 h as reported by (Pakostova,

D7.2 |

2017) in the framework of the BIOMOre project would not result in a more economic operation compared to H2O2 dosage. Based on first estimates, HRT should be in the order of only 1 h, subject to final testing and verification of cost figures based on a long-term indistrial-scale trial.

3.1.3. Oxidative vs. non-oxidative ISR of Cu

It has been demonstrated in Section 3.1.2 that the OPEX for oxidizing the lixiviant are predominant in the case of ISR applied to reduced (sulfidic) Cu ore (simulated for chal-cocite without redox competitors). Assuming the leaching of an oxidized ore, the costs for Fe2(SO4)3 and oxidation of lixiviant do no longer apply, thus, reducing the specific costs from 2.81 €/kg to 0.70 €/kg, see specific distribution of costs in Figure 15.

Figure 15: As for Figure 14, but assuming non-oxidative Cu ISR (i.e. leaching from oxidized

Cu minerals). Total specific OPEX: 1.09 €/kg.

The leaching rate for the non-oxidative leaching has been chosen to be the same as for the oxidative leaching (hypothetically).

3.1.4. Dependence of costs on the depth of the deposit

The depth of the deposit determines the costs for drilling, well construction and pump-ing (extraction pumps) mainly. A reduction of depth from 250 m (as assumed in Section 3.1.2) to 50 m reduces both CAPEX (from 0.48 €/kg to 0.14 €/kg) and OPEX (from 2.81 €/kg to 2.70 €/kg), thus, resulting in total costs of 2.84 €/kg for pure ISR and 4.30 €/kg for producing the pure metal. The CAPEX cost structure is shown in Figure 16. The well construction costs are still dominant (now 57 % instead of 79 % for 250 m depth).

D7.2 |

Figure 16: As for Figure 13, but assuming a deposit depth of 50 m only (instead of 250 m).

Total specific CAPEX: 0.14 €/kg.

Vice versa, the CAPEX and OPEX increase significantly as the depth of the deposit increase. For 1,000 m, the CAPEX increase from 0.48 €/kg to 1.77 €/kg, whereas the OPEX increase from 2.81 €/kg to 3.21 €/kg (total ISR costs: 4.98 €/kg; total Cu pro-duction costs: 6.44 €/kg) with reference to the 250 m depth case.

Note that this simulation has been performed for ISR favourable conditions (3 wt% ore grade at 25 % porosity and neglecting redox competitors).

3.1.5. Systematics of factors influencing costs and economic feasibility

The feasibility criteria of ISR indicate a complex influence on the final costs. In order to systematize the dependence of costs on individual parameters, the “standard” sce-nario described in Section 3.1.2 has been used as a reference. The parameter varia-tion has involved some optimization procedures (i.e. adjustment of operational param-eters including flow rate, cut-off concentration and others). The calculated specific costs for ISR (without metallurgical processing) were normalized to the specific costs of the standard scenario for comparability (i.e. the cost ratio has been deduced to demonstrate the relative influence). Note that the trends change quantitatively for dif-ferent reference scenarios, but the qualitative dependences are about the same. The identified critical parameters affecting the economic feasibility of acidic ISR include:

Calcite (carbonate) abundance. Ore grade. Depth of or body. Permeability (related to porosity). Spacing between injection and extraction wells. Oxidant consumption by the target mineral and competitors. Soluble chloride abundance in case of microbial processing (bio-leaching); qual-

itative discussion only.

D7.2 |

Figure 17 shows the significant increase in costs with increasing calcite abundance. Initially, the increase in costs is almost constant, reflecting the increasing acid con-sumption and leaching time due to the retarded breakthrough of acid (deliverable D7.1). This is valid for all carbonates, e.g. dolomite, siderite. With calcite abundances ≥ 1 wt%, the costs increase becomes non-linear due to well repair and maintenance efforts in view of the precipitation of gypsum, which clogs pores, screens, and process equipment. Thus, calcite abundances up to about 1 wt% could be tolerable, but acidic ISR is no longer feasible at abundances in excess of 1.5 wt%. This study explicitly performed for calcite applies to other calcareous minerals accordingly.

Figure 17: Normalized specific ISR costs (CAPEX + OPEX) in dependence on the calcite

abundance in the ore deposit (red line: reference case at 0 wt% calcite)

Figure 18 represents the normalized specific costs in dependence on the ore grade.

Figure 18: Normalized specific ISR costs (CAPEX + OPEX) in dependence on ore grade

(red line: reference case at 1 wt% copper)

D7.2 |

The higher the ore grade the higher the Cu concentration in the PLS, provided that sufficient oxidant (ferric iron) is injected into the system to avoid or minimize thermo-dynamically constrained leaching. The cut-off concentration (metal concentration in the PLS as criterion for wellfield shut-down) has to be adjusted. The trend curve in Figure 18 has been deduced from the KiLea-Hy model linked to the economic module by accordingly adjusting operational parameter for optimization. In this regard, it is implicit to some extent.

As already discussed in Section 3.1.4, the total ISR costs depend on the depth of the deposit significantly. The deeper the deposit the higher the well construction costs (CAPEX) and the higher the costs for electrical energy to pump the PLS to the surface (OPEX). Drilling costs are dependent on the rock types that need to be drilled through. With increasing depth, the probability of complications during drilling increases; this is indicated by the dashed line in Figure 19.

Figure 19: Normalized specific ISR costs (CAPEX + OPEX) in dependence on the depth of

the ore body (red line: reference case at 250 m depth)

The costs figures as function of depth is mainly determined by the drilling technique and equipment used. More advanced drilling methods may change the figure to some extent. Alternatively, wellfields with horizontal wells may be less cost-intensive due to the reduced number of wells required. However, there is only experience with horizon-tal wells regarding heat transport in geothermal applications rather than with mass transport applications required for leaching. Further investigations are required to ac-cess the technical and economic feasibility of horizontal wellfields for ISR. Definitely, it would only be applicable to consolidated rock formations (i.e. stable enough to con-struct the wells including screens without risking drill holes to collapse).

The dependence of the specific ISR costs on the permeability of the ore body is pre-sented in Figure 20. The relationship between permeability and (effective, i.e. free-fluid) porosity can be expressed by the Kozeny-Carman equation, valid for sedimen-tary deposits (cf. deliverable D7.1). Figure 20 includes the approximate porosity

D7.2 |

windows corresponding to the permeability scale for sediments characterized by a me-dium average grain size and low abundance of silt/clay and/or cementing minerals.

Figure 20: Normalized specific ISR costs (CAPEX + OPEX) in dependence on permeability corresponding to about the porosity figures included (red line: reference case 20 % porosity

or 0.6 m/d permeability)

The trend curve indicates that the specific ISR production costs increase at low poros-ity/permeability. Under such conditions the effective reaction space (pore volume and inner reactive surface) for ISR becomes limited and the flow rate would be too low for a quantitative production rate (cf. deliverables D2.4 and D7.1). At a permeability of less than 0.10 m/d (about 120 mD) the wellfield-based in-situ recovery is no longer economic in most application cases due to the very low flow rates achievable (i.e. not sufficient for transporting enough reactant into the leaching system and to get the leached metal out of it).

Fracturing (in particular, hydro-fracturing) could be applied to increase the permeability by creating cracks/fissures. Due to the incompressibility of rock the effect of fracturing is limited. Whereas the permeability can be increased by several 10 percent, the effect on the porosity increase is marginal.

Figure 21 represents the specific ISR costs in dependence on the spacing between injection and extraction wells. Two application cases were considered: high-grade ore and low-grade ore.

There is an overlap of two main dependences:

The lower the spacing the higher the drilling and well construction costs per kg metal (spacing determines the ore body volume per pattern and, hence, the amount of metal to be leached in dependence on the ore grade).

The higher the spacing the larger the relative volume of dead zones with insuf-ficient or no contact of the lixiviant to the ore. This is illustrated in the lower part

D7.2 |

of Figure 21. In addition, the dilution of the PLS by groundwater that is sucked in by the extractor increases (a function of leaching time).

The ore grade influences the relative weight of both effects.

Figure 21: Normalized specific ISR costs (upper diagram), distinguished for a low-grade and

high-grade deposit, and ore zone available for the leach solution at 50 days after wellfield start-up (lower figure) dependent on spacing between injection and extraction wells in the

wellfield.

Finally, an important cost factor applies to the ISR of reduced minerals. The higher the number of electrons ne that needs to be exchanged to mobilize (leach) just one metal atom (cf. Figure 22), the higher the consumption of chemicals to re-oxidise ferrous to ferric iron or the higher the costs to run a biooxidation plant (cf. deliverable D7.1). In addition to a typical base load to oxidise the lixiviant, the costs for re-oxidation are proportional to ne (at a gradient that is influenced by competing reductants in the ore formation).

D7.2 |

Figure 22: Electron exchange per metal dissolved dependent on the mineralization.

[Note: The chemical structure of covellite CuS is similar to sphalerite ZnS.]

Despite the target mineral, competitive minerals for the oxidant have to be considered in the oxidant consumption costs. Simultaneous dissolution of redox minerals against the thermodynamic order are often observed due to probable differences in:

Availability of minerals to the leachant in microscale and along the flow path.

Reaction surface of the minerals.

Slower chemical reaction kinetics compared to other minerals, etc.

While some of the leached minerals may add revenue, if a metal can be purified eco-nomically, e.g. Zn from sphalerite, or if a chemical reagent can be recycled in the pro-cess, e.g. sulfuric acid, others produce costs related to water treatment and waste management (e.g. Fe from pyrite).

While the in-situ conditions and therewith the production rate vary in space and time, the purification process is most economic if copper concentrations in the PLS are nearly constant over time. Therefore, the wellfield operation requires both a technical operation and a wellfield production management (in particular, by realizing a consec-utive wellfield setup) to ensure an almost constant production rate.

D7.2 |

3.2. Cost estimates for copper purification

Deliverable D4.1 compares the relative incremental costs of production with the reve-nue of different metals and intermediate products for a range of suitable recovery unit operations using a techno-economic screening process. The following processes may be economically applicable for the BIOMOre test conditions:

Copper by SX and EW.

Copper by IX and EW.

Copper sulfide precipitation.

The recovery of additional metals as by-products in the BIOMOre process is not con-sidered in the present CAPEX/OPEX analysis, but is definitely an option for future potential industrial applications.

3.2.1. Cost effective flow sheet

CLC has identified and analysed the most relevant process stages (and associated costs) required to extract copper by means of an in-situ (bio-)leaching process. In order to perform CAPEX and OPEX estimations for the BIOMOre process, CLC provides a list with the most important processing steps:

Water & Acid washing: Due to the high concentration of chlorides in the BIO-MOre pilot test ore, a new pre-treatment consisting of water plus acid washing has been studied and included in the design. This eliminates a large amount of chlorides that prevent microorganisms from coexisting.

In-situ (bio-)leaching of copper ores: Three stage counter-current cascade with integrated ferrous ion bio-oxidation, adapted to mining exploitation plan and phi-losophy defined by the operators (KGHM).

Copper purification by Solvent Extraction or Ion Exchange technologies.

Copper production by EW technology.

Effluent Treatment Plant (ETP): waste water treatment.

The mass balance model METSIM is used to simulate these processing steps. The basis of the mass balance model is a process flow diagram. Based on its broad know-how and experience in copper extraction from sulfide ores after six years of successful operation of its Hydrometallurgical Plant in Seville, Spain, CLC has undertaken a de-tailed analysis of some industrial bio- and hydrometallurgical technologies that may be suitable for the BIOMOre process. The main output of such studies is that copper extraction in the BIOMOre process is expected to be similar to copper (bio) heap leach-ing operations (cf. Figure 23).

D7.2 |

Figure 23: Proposed flow sheet for the BIOMOre process

NOTE: (*) Groundwater is incorporated into the process water flows as it must be considered in the overall water balance.

The key of this technological approach is to divide the mining exploitation area into blocks, to be similar to heap stockpiles. Then, to follow a proper operating plan based on counter-current leaching associated with a sequential bio-oxidation process from low-grade to high-grade copper content in solution:

High grade (new block).

Medium grade (mid-block that is already under exploitation).

Low grade (final block, end or exhausted).

D7.2 |

3.2.2. Main streams and effluent compositions

The following process design conditions, after discussion and agreement with the part-ners involved in this task, are considered as a basis for the CAPEX/OPEX estimates using METSIM software:

Production of copper cathodes: 20,000 t/y.

Independent variable: PLS = 2 to 5 g Cu/L.

Total iron consumption: 20% over stoichiometric (calculated on Cu basis).

Acid content in injected solutions: 20 g/l.

Cu recovered in overall leaching process: 20%.

The defined design conditions, in combination with deposit-relevant data, allow the derivation of base parameters for material balance. As pointed out in Section 3.1, the ferric iron concentration in the leachant has to be adjusted to realize the target Cu concentration in the PLS in dependence on Cu sulfide mineralogy to meet redox (elec-tron-exchange) conditions.

Production:

Assuming an annual production of copper cathodes of 20,000 t/y, and an operation time of 350 d/y, we obtain:

24 h/d ∙ 350 d/y = 8400 h/y.

The resulting copper extraction obtained is:

(20,000 t/y Cu / 8400 h/y) ∙ 1000 = 2,381 kg Cu/h.

Average sandstone layer composition:

Data on the composition of the KGHM deposit’s sandstone layer was provided by IMN according to Monografia KGHM, Lubin 2007, p. 198.

Components considered in the database of the METSIM model are presented in Table 8. Minor components with very low concentrations have been neglected in the model, i.e. Ag (29 ppm), Ni (46ppm), Co (19 ppm), V (59 ppm), and Mo (40 ppm). Using a copper grade of 2.67 % (from Table 8) as an average, the following (quite conserva-tive) calculations are performed:

Amount of ore treated: 2,381 kg Cu/h / (2.67 % 20%) = 445,871 kg/h.

Amount of copper in the ore body: 445,871 kg/h 2.67 % = 11,905 kg Cu/h.

Residual copper (amount of unleached copper): 11,904.76 kg Cu/h – 2,381 kg Cu/h = 9,523.81 kg Cu/h.

D7.2 |

Table 8: KGHM deposit’s sandstone layer.

Component Composition (%) Component Composition (%)

SiO3 69.8 FeS2 0.19

Al2O3 4.22 FeO 0.62

CO2 6.5 Fe2O3 0.53

CaO 7.34 Cu 2.67

MgO 2.19 Zn 0.04

Ss 0.82 Pb 0.05

Average mineralogical composition of sandstone layer:

Table 9 summarizes the data (according to IMN’s data) that have been taken into ac-count in the mass balance of METSIM.

Table 9: Sandstone layer composition in the Rudna mine.

Component Composition (%)

Clay minerals 14.8

Quartz 71.4

Dolomite 5.1

Calcite 4.8

Cu sulfide minerals 3

Composition of copper sulfide minerals in sandstone layer:

Table 10 summarizes the different mineral species of copper in the Rudna mine (ac-cording to IMN’s data) and the used adjustment for the MetSim model.

Using an assumed average copper grade of 2.67%, the composition of copper sulfide minerals in the sandstone would be:

Chalcocite: 2.67% (70.61% / 100) = 1.89% MCu2S/MCu2 = 2.36%.

Digenite: 2.67% (6.27% / 100) = 0.17% MCu9S5/MCu9 = 0.21%.

Bornite: 2.67% (9.44% / 100) = 0.25% MCu5FeS4/MCu5 = 0.40%.

D7.2 |

Chalcopyrite: 2.67% (4.82% / 100) = 0.13% MCuFeS2/MCu = 0.37%.

Covellite: 2.67% (8.86% / 100) = 0.24% MCuS/MCu = 0.36%.

Table 10: Copper minerals of the sandstone layer in the Rudna mine.

Component Minimum Maximum Average Adjustment

Chalcocite 52.5 94.6 69.6 70.61 (*)

Digenite 2.1 11.5 6.27 6.27

Bornite 2.1 23.8 9.44 9.44

Chalcopyrite 0.13 13.5 4.82 4.82

Covellite 0.82 15.0 8.86 8.86

Total 98.99 100.0

(*) Chalcocite, the quantitatively main copper sulfide, was adjusted to add up to a total of 100%.

3.2.3. Mass balances

The in-situ bioleaching modelled in METSIM follows the flow sheet shown in Figure 23. Three blocks are considered to be leached in a counter-current regime. The block stages over time are: first leached mineral block (new block 3), second leached block, if a significant portion of the copper to be recovered is leached (mid-block 2), and, ultimately, the final leached block, if the targeted copper recovery is almost achieved (final block 1). In contrast, the leach solution passes the blocks from low to high copper grade (blocks 1 to 3).

PLS estimates:

Initially, two levels of copper concentrations in the PLS have been considered:

Low: 2.381 kg Cu/h / 2.5 g/L = 952 m³/h.

High: 2.381 kg Cu/h / 5.0 g/L = 476 m³/h.

During the development of the material balance, the scenario for 2.5 g/L of copper has been selected as most reasonable. This is due to the high dolomite and calcite content of the ore that causes calcium saturation during the process, thus, impairing flow cir-culation within the ore block.

D7.2 |

Leaching recovery:

For the mass balance, a conservative total copper leaching recovery of 20% of from a block is assumed. This percentage was divided into each of the blocks according to their status (cf. Table 11).

Table 11: Copper recovery rate from the blocks during ferric leaching.

Block status Leaching recovery

New block 12%

Mid-block 6%

Final block 2%

Table 12 presents the leaching recoveries with respect to every copper mineral spe-cies within each block.

Table 12: Recovery of individual mineral species during the ore leaching.

Prediction for leaching of species in the ore leaching

Species Chalcocite Digenite Bornite Chalcopyrite Covellite

Block es-timation

13% 9% 9% 3% 13%

7% 5% 5% 2% 6%

2% 2% 2% 1% 2%

Total 22% 15% 15% 5% 21%

Figure 24 to Figure 28 present the exemplarily mass balances from METSIM for the 5 main processing steps defined in Section 3.2.1. The basic process flow diagrams of the main processes are presented in Annex 7.2. Outputs of such a mass balance were used to estimate the CAPEX and OPEX including the items:

Reagent consumption and associated costs.

Personnel to operate the process plant and corresponding costs.

Maintenance costs for the process plant.

Overall OPEX.

D7.2 |

Figure 24: Mass balance from METSIM for water and acid washing.

Figure 25: Mass balance from METSIM for bio-leaching and bio-oxidation.

D7.2 |

Figure 26: Mass balance from METSIM for solvent extraction.

Figure 27: Mass balance from METSIM for copper electrowinning.

D7.2 |

Figure 28: Mass balance from METSIM for Effluent Treatment Plant (ETP).

D7.2 |

3.2.4. CAPEX and OPEX estimates Copper recovery from PLS

The capital and operating costs for processing the PLS have been calculated, taking into account two purification processes:

Solvent extraction (SX) followed by electrowinning, including ETP.

Ion exchange (IX) followed by electrowinning, including ETP.

Both purification processes yield the same product: copper cathodes. Figure 29 rep-resents the scheme for the recovery of copper from the PLS obtained by SX.

Figure 29: SX- EW flow sheet.

The CAPEX figures of the purification process using SX are presented in Table 13. The final estimate based on the PFD and the mass balance modelled with METSIM is 84.71 M €, equivalent to 4,235 € per ton copper.

Table 13: Capital costs of the purification process using solvent extraction.

Solvent Extraction

Electrowinning

PLS RE Cuº

Raffinate

Cost of main equipment by area SX ‐ EW ‐ EPT

Unit Value

Solvent Extraction M € 11,95

Electrowining M € 23,91

Efluent Treatment Plant (ETP) M € 2,26

Main Equipment Total M € 38,12

Main Process Equipment 45% M € 38,12

Project Cots (engineering + process mgmt. + site work) 10% M € 8,47

Piping 9% M € 7,62

Process Electrification 6% M € 5,08

Process automation, instrumentation & control 5% M € 4,2

Installation, construction, building, HVAC* 25% M € 21,2

Total Investment Cost 100% M € 84,7

*HVAC (Heating, Ventilation and Air Condition)

Total CAPEX 84,71 M €

€/ton Cu 4235

CAPEX

D7.2 |

The OPEX estimates include all stages of the purification process using SX (cf. Table 14). Annual costs of 50.11 M € have been calculated, corresponding to 2,506 € per ton copper.

Table 14: Operating costs of the purification process using solvent extraction.

The second purification process option is to use IX instead of SX for the copper recov-ery from the PLS. The advantage of IX is that no organic media, which could interfere with the bio-oxidation and/or leaching process, is required. Furthermore, the potential contamination of the leaching field by an organic medium could be a regulatory issue. Figure 30 represents the IX-based purification scheme, where the SX process is re-placed by the IX process.

Figure 30: IX – EW flowsheet.

CAPEX and OPEX of the IX purification process are presented in

Unit Value

Reagents M € 36,17

Utilities M € 6,954

Labour (maintenance labour included) M € 4,45

Maintenance M € 2,54

Total OPEX 50,11 M €

€/ton Cu 2506

OPEX

Ion Exchange Electrowinning

PLS RE Cuº

Raffinate

D7.2 |

Table 15 and Table 16. Compared to the first purification option using SX, the CAPEX are significantly higher at 103.21 M €, corresponding to 5,160 € per ton copper. The OPEX of both purification options are in the same range, with annual costs of 50.75 M €, corresponding to 2,538 € per ton copper for option 2.

In summary, the purification process using solvent extraction is economically favoured. Thus, if the discussed technological issues of SX are non-restrictive for the process, SX is the recommended technology for copper purification and concentration in the BIOMOre process.

D7.2 |

Table 15: Capital costs of the purification process using ion exchange.

Table 16: Operating costs of the purification process using ion exchange.

Finally, the costs have to be related to the value of the product(s) to determine the economic viability of a technological process. In Task 3 of Work Package 7 (cf. deliv-erable D7.3), an economic evaluation of a potential in-situ bioleaching operation at industrial scale is performed based on these cost estimates. There, the copper price development and copper mining costs related to the price are reviewed.

Cost of main equipment by area IX ‐ EW ‐ EPT

Unit Value

Ion Exchange M € 20,27

Electrowining M € 23,91

Efluent Treatment Plant (ETP) M € 2,26

Main Equipment Total M € 46,44

Main Process Equipment 45% M € 46,43

Project Cots (engineering + process mgmt. + site wo 10% M € 10,32

Piping 9% M € 9,29

Process Electrification 6% M € 6,19

Process automation, instrumentation & control 5% M € 5,2

Installation, construction, building, HVAC* 25% M € 25,8

Total Investment Cost M € 103,2

*HVAC (Heating, Ventilation and Air Condition)

Total CAPEX 103,21 M €

€/ton Cu 5160

CAPEX

Unit Value

Reagents M € 36,25

Utilities M € 6,95

Labour (maintenance labour included) M € 4,45

Maintenance M € 3,10

Total OPEX 50,75 M €

€/ton Cu 2538

OPEX

D7.2 |

4. Effects on copper prices

Common factors such as the copper grade and purity as well as the presence of im-purity elements such as chlorides and sulfates will affect the price generated by all forms of the product (cf. Table 17). The selection of the site location will also have a significant effect on a number of factors, including transport options and associated costs.

Table 17: Copper products and their value from different small-scale production processes.

Production Process Product Value $/t Contained copper

Crystallization Copper Sulfate $ 7300

SX/EW Copper Cathode $ 6600

Cementation Copper Cement $ 5600

Sulfide Precipitation Copper Sulfide $ 5000

The copper price in € is presented in Figure 31. Compared to the obtained purification operating costs, they show promising results.

Figure 31: Copper price over the last few years.

Over the past five years, the industry has experienced an increase in both the capital and operating costs, following this thirty year trend. Figure 32 illustrates the copper price in real 2009$ terms over the past century with a schematic for future prices. It shows that the copper price is cyclical and from the 1970s to mind-2000s, it appeared to be in structural decline.

D7.2 |

Figure 32: Long term copper prices.

The Wood Mackenzie approach for the longer-term forecast of copper prices – in this context for the next price cycle of approximately ten years – is driven by our estimates on the demand for refined copper.

These estimates are based on the intensity of copper use in various sectors and re-gions, and the growth in end-use demand. With global industrial production, the key economic driver, the growth is assumed to remain cyclical. The impact of global de-mographic changes and the urbanisation of economies, for example in China, are em-bedded in the modelling of demand.

Scrap usage:

In modelling the potential demand for refined copper, the proportion of the demand that will be met by processing scraps must be estimated. The demand itself is impacted by:

Advances in technology.

Changes in environmental legislation.

Economics of scrap processing. In 2009, scraps processed at smelters and refineries accounted for 16 % of the re-fined production at 2.9 Mt (cf.

D7.2 |

Table 18).

D7.2 |

Table 18: Scrap contribution as percentage of total use of copper.

The cost price relationship:

When the mine supply exceeds the demand, prices typically fall to a level that places a sufficient pressure on mines that causes them to put a number of mines under care and maintenance – or to permanently close. These prices are directly related to the expected production cost of the industry. This relationship is the best indicator of prices in the future, the cost of investing in new capacity and the financial returns that these investments can potentially generate.

Figure 33: The cost price relationship in 2009$.

Figure 33 represents the Composite Cash Cost versus the time in years. Composite Cash Cost is a company's cost to finance its business, determined by, and also re-ferred to as "weighted average cost of capital" or WACC. The calculation involves mul-tiplying the cost of each capital component by its proportional weight and taking the sum of the results. The correlation between industry costs and the copper price is less strong over the entire cycle than during periods of excess supply.

D7.2 |

How capital costs have escalated:

Until around 2005, it was difficult to discern trends in capital costs when historic costs were rebased into current real terms. From 2005 to 2008, however, capital costs in-creased considerably. This may potentially shift the incentive price curve higher, as-suming costs stay at 2008 levels. Figure 34 and Figure 35 show how the capital costs of individual projects have increased over time, related to the initial level of capital costs forecast.

Figure 34: Mines commissioned in 2006 and 2007.

Figure 35: Mines commissioned in 2008 and 2009.

D7.2 |

The average total cost escalation for a group of six mines commissioned over 2006 and 2007 in the USA was 50%, or around 20% per year, over the period from the initial capital cost estimate until completion (cf. Figure 34). The total increase in the capital costs forecast for a representative group of seven mines commissioned over 2008 and 2009 averages 120%, compared with the initial estimate, or around 40% of the capital cost increases per year over their construction period (cf. Figure 35).

D7.2 |

5. Cu ISR cost review and conclusions

This report describes the CAPEX/OPEX analysis and assessment of the BIOMOre process, clearly distinguishing between the actual ISR technology (in-situ leaching to enrich the metal of interest in the pregnant leach solution PLS) and the subsequent metallurgical processing, i.e. separation of the metal out of the PLS by either SX or IX followed by the refining to a pure metal (electrowinning).

Regarding the primary recovery (ISR), the effect of both deposit-dependent and oper-ational parameters on the specific cost figures (€ per kg metal) of an ISR process was quantified with reference to economic criteria. The production rates simulated by the generic model software KiLea-Hy were combined with an economic model to calculate CAPEX/OPEX for a quantitative assessment of an acid ISR operation applied to re-duced (sulfidic) ore. These combination allows the quantification of specific ISR costs (in €/kg metal) dependent on deposit-related and operational parameters. The main conclusions from this study include:

Too high abundances of calcareous minerals constrain the applicability of acid ISR. In addition to the economic effect of increased costs to acidify the ore for-mation, increased repair & maintenance costs apply in view of quantitative gyp-sum precipitation (in the formation itself clogging pores, in the screens as well as in pipes and plant components). Whereas calcite concentrations (reference mineral) in the order 1 wt% become critical, ISR is no longer applicable at abun-dances in excess of 1.5-2.0 wt% due to excessive costs and constrained tech-nical feasibility.

High ore grades are favourable obviously. Metals of high value (market price) can be recovered from low-grade ores (in the order of 0.1 wt% or even less) economically. The economically feasible ore grade limit is strongly correlated with the price level for the metal(s) of interest. In the case of Cu, ISR might become noneconomic at ore grades less than 0.2-0.5 wt% (also depending on other deposit parameters).

ISR CAPEX costs are dominated by wellfield construction costs including drill-ing. Again the economically determined maximum depth depends on the market value of the metal. Whereas depths up to 1,000 m are feasible for high-value metals, Cu ISR is limited to about 200 m depth, if considering conventional drill-ing and construction of vertical injection and extraction wells. However, innova-tive drilling and well construction technologies, optionally considering horizontal wells (only applicable to consolidated rock formations), could increase this limit substantially.

KiLea-Hy simulations combined with estimates of CAPEX and OPEX have demonstrated the influence of porosity and permeability . Both porosity and permeability are interrelated, however, depending on lithological conditions. Based on the simulations, Cu ISR is feasible in the window > 10 % and per-meability > 0.1 m/d (120 mD). For higher price metals, this window could ex-tend to lower values. Fracturing (cf. D7.1 for applicable technologies) could en-hance the permeability in certain limits (several 10 %), but an increase of

D7.2 |

porosity is almost impossible due to the incompressibility of rock. Whereas per-meability determines the pore volume exchange rate (in other words, the solid/liquid ratio in time), the porosity (distribution) defines the available reaction space and leachant/ore interface (inner reactive surface). Both factors deter-mine the final production rate and, hence, cost figures, considerably.

Spacing (between injection and extraction wells) is a key parameter for wellfield design. Whereas low spacing increases the wellfield construction costs, larger spacing decreases the leaching efficiency (recovery) due to increasing dead zones (without sufficient or no contact of leachant to ore) and dilution effects. The optimum (usually in the order of 30-50 m) is, however, dependent on the ore grade and the value of the metal. KiLea-Hy is an efficient tool for the opti-mization of the wellfield design. Fine-tuned optimization is based on 3D reactive transport modelling (state-of-the-art approach).

Whereas it could be demonstrated that the ISR of Cu from oxidised ore bodies results in total specific costs in the order or less than the specific market price, the application of ISR (in particular, the BIOMOre technology including the ex-situ biooxidation of iron as a direct catalyser of leaching) to reduced (sulfidic) Cu ores increases the costs due to the efforts to oxidise the leachant. KiLea-Hy simulations have shown that the costs are dependent on (i) the number of elec-trons to be exchanged for leaching one atom of Cu (dependent on the mineral form of the sulfidic Cu ore) and (ii) the impact of reducing competitors consum-ing oxidant as well (dependent on the abundance of such mineral phases/or-ganics and specific reaction kinetics).

The above-described main influences of the most significant parameters on specific ISR costs have been systematized for typical deposit and ISR operation conditions.

Additionally, the cost figures have been completed by the detailed economic assess-ment of the subsequent metallurgical processing. CLC applied the software tool METSIM to the complete BIOMOre processing (IX/SX followed by electrowinning) con-sidering: (i) water & acid washing, (ii) leaching & bio-oxidation, (iii) Cu extraction from PLS, (iv) Cu electrowinning, and (v) effluent treatment. The process flow diagrams are the basis for the CAPEX/OPEX estimates of the BIOMOre process including an eco-nomic comparison between metal recovery by SX and IX.

In a next step (sufficient production rate implied) the ISR operation costs are combined with the subsequent hydrometallurgical processing. Therefore, CLC extended the pro-cess flow model in METSIM to simulate the complete BIOMOre process inclusive the new stages of water and acid washing applied in the Rudna Mine block reactor. The stages were required to avoid (i) high levels of chloride feeding the bacterial culture, and (ii) precipitation of hydrous ferric oxides in the main leaching stage.

Using above mentioned information the Material Balance for the process has been developed producing enough information to estimate process reagents and utilities consumptions to be integrated in the operating expenditures (OPEX). Labour and

D7.2 |

maintenance cost have been estimated using average data from similar industrial ap-plications to finally calculate the complete BIOMOre OPEX for the process plant.

With the METSIM model as reference, the total investment cost has been developed by means of the conceptual engineering study of each process stage. CAPEX esti-mates have been developed using stochastic estimating methods such as cost/capac-ity curves and factors, scale of operations factors, Lang factors, Hand factors, Chilton factors, Peters-Timmerhaus factors, Guthrie factors, and other parametric and model-ling techniques. The comparison showed that SX is economically advantageous.

D7.2 |

6. References

Amrich Minerals, L., 2013. Santa Cruz Project, Pinal County, Arizona.

AACE® International Recommended Practice No. 18R-97: Cost estimate classification system – as applied in engineering, procurement, and construction for the process industries.

Amrich Minerals, L., 2013. Santa Cruz Project, Pinal County, Arizona.

Briggs, D.F., Allison, M.L., 2015. Recovery of Copper by Solution Mining Methods, Contributed Report CR-15-A.

Carlson, N., Zonge, K., Ring, G., Rex, M., 2000. Fluid Flow Mapping at a Copper Leaching Operation in Arizona. Lead. Edge 19.

CLC Internal production costs database, 2017.

CLC Internal equipment investment cost database, 2017.

CLC internal proposal from Johnson Matthey for IX Cu recovery from solution, 2017.

Coombs, D., Hunt, B. (a Wood Mackenzie Company) (2010) The Influence of Costs on Future Copper Prices, International Cooper Conference 2010, Hamburg, Germany.

Crundwell, F. (2006) Process and Economic Considerations in Copper Metallurgy, CM Solutions (Pty) Ltd, Presentation.

CSIRO, 2015. More from less GETTING THE MOST FROM AUSTRALIAN ORES. Resourceful.

Fugleberg, S. (1999) Finnish expert report on best available techniques in zinc production, In series: The Finnish Environment 315. https://helda.helsinki.fi/handle/10138/43123.

Hackman, D.B., 1982. The evaluation of supergene copper deposits for in situ leaching 125.

Kennedy, M.W. (2014) Metallurgical plant optimization through the use of flowsheet simulation modelling, Symposium on Pyrometallurgy, The Minerals, Metals & Materials Society.

Meka, L. (2001) Copper SX-EW Plant design. A case study of a low cost approach, AUSEMCO Ltd., https://www.ausenco.com/.

Norgate, T.E., Rankin, W.J. (2002) An Environmental Assessment of Lead and Zinc Production Processes, Green Processing 2002 - Proceedings: International Conference on the Sustainable Processing of Minerals, pp. 177-184.

D7.2 |

Pakostova, E., Grail, B.M., Johnson, D.B., 2017. Indirect oxidative bioleaching of a polymetallic black schist sulfide ore 106, 102–107. https://doi.org/10.1016/j.mineng.2016.08.028

Pugliese, J.M., 1989. Generic design manual: cost model for in situ copper mining, in: In Situ Leach Mining. p. 107.

Qiuyue, Z., Tingan, Z., Xiaofeng, Z. (2013) Application of Process Simulation Software METSIM in Metallurgy, iaes, Vol 10, No 8.

Rosenbaum, J.B., McKinney, W.A., 1970. In situ recovery of copper from sulfide ore bodies following nuclear fracturing.

Sakata, M., Imai, T., Horiuchi, A., Morita, K., Kondo, Y., Kusano, M. (2005) Zinc Smelters Survey, In International symposium on lead and zinc processing; pp. 139-202, International symposium on lead and zinc processing, Kyoto, Japan.

Sarangi, A. K., & Beri, K. K. (2000). Uranium mining by in-situ leaching. In Proceedings of the International converence on Technology management for mining, processing and environment.

Seredkin, M., Zabolotsky, A., Jeffress, G., 2016. In situ recovery, an alternative to conventional methods of mining: Exploration, resource estimation, environmental issues, project evaluation and economics. Ore Geol. Rev. 79, 500–515. https://doi.org/10.1016/j.oregeorev.2016.06.016

Sinclair, R.J. (2005) The Extractive Metallurgy of Zinc. In Spectrum Series, AusIMM.

Sinclair, L., Thompson, J., 2015. In situ leaching of copper: Challenges and future prospects. Hydrometallurgy 157, 306–324. https://doi.org/10.1016/j.hydromet.2015.08.022

Taylor, G., Farrington, V., Woods, P., Ring, R., & Molloy, R. (2004). Review of Envi-ronmental Impacts of the Acid In-situ Leach Uranium Mining Process. CSIRO Land and Water Client Report.

D7.2 |

7. Annex

7.1. Glossary of terms

[In the style of (ref. IAEA ISL), reworked and extended.]

Term Definition, explanation

Aquifer A permeable underground sediment or rock formation capa-ble of storing and allowing the flow of water (groundwater).

Aquitard A low-permeability underground sediment or rock formation that retards but may not entirely prevent the flow of water (groundwater).

Baseline Data acquired to identify the state of the environment prior to any disturbance from mining. It should give a pre-mining in-ventory of factors such as the diversity and abundance of flora and fauna, agricultural or pastoral activities and produc-tivities, and quality of air and water, particularly groundwater. The values acquired can be used as a benchmark for final mine rehabilitation.

Social baseline data is also relevant to most projects.

Bleed In most but not necessarily all ISL operations, ‘bleed’ is a slight excess of pumped water compared to injected water is maintained, which may vary from <1% to a few % of the wa-ter extracted. This helps maintain mining solutions within the planned mined areas, but also generates excess water that must be used or disposed of.

Block Leaching The leaching of a block of ore accessed by a conventional underground mine, where mining solution is passed through a block of blasted ore and recovered. Whilst this has similari-ties to ISL mining as described here, it is not included except by passing reference.

Confinement (In ISR) The (vertical) hydraulic separation of the (intended) ISR mining zone (wellfield) in an aquifer by adjacent aqui-tards. In general, it requires the absence of unwanted (hy-draulic) connectivity, e.g. by tectonic faults.

Cut-off grade Metal concentration in the lixiviant of a depleting wellfield, at which the ISR operation becomes uneconomic. Shutdown criterion, usually applied to individual extraction wells.

D7.2 |

Effluent Wastewater from chemical/metallurgical processing

EIA/EIS Environmental Impact Assessment/ Statement, an assess-ment or statement of the environmental impacts of a project, usually incorporating some form of risk assessment and measures proposed to minimize identified, expected or pos-sible negative impacts. May include a social impact assess-ment.

Elution IX process. Recovery of uranium (in form of anionic uranyl complexes) from the loaded resin beads by a highly ionic el-uant stream (containing nitrate, chloride or others), thus, forming a pregnant eluate with a high uranium concentration for further processing. To be followed by a resin regenera-tion (also referred to as conversion) for reuse in the sorption circuit (c.f. sorption).

Excursion Movement of mining or waste solution beyond the area in-tended/permitted.

Extraction well A cased well with a filter section through which mining solu-tion is removed from the orebody being mined by pumping.

GT Grade x Thickness (of an orebody) in wt%m, to be distin-guished from Productivity.

Hydraulic conduc-tivity (permeabil-ity)

The inherent ability of porous rock to transmit fluid (c.f. trans-missivity)

Injection well A cased well with filter section through which mining solution is introduced to the orebody being mined, either pumped or under gravity. The mining solution is removed via an extrac-tion (or recovery) well.

ISL/ISR In-situ leaching / In-situ recovery, a form of solution mining where mining solution is circulated underground through an sedimentary, saturated ore body to dissolve and bring the target material to the surface for further recovery.

IX (Ion Exchange) Transfer (absorption/desorption) of ions between a solution and ion exchange sites on surfaces (on and inside porous IX resin beads in uranium recovery applications). Together with SX, a predominant technology to recover a metal from loaded mining solutions in ISL applications (cf. sorption and elution).

D7.2 |

Lixiviant Synonym for mining solution in ISR/ISL applications. Often referred to as Lix.

Further synonyms:

Pregnant lix(iviant): leachate, pregnant leaching solution (PLS), production solution.

Barren lix(iviant): mining solution after metal extraction by IX or SX, injection solution (after refortification).

Mining Solution Usually local groundwater that is chemically modified to cause it to dissolve the target mineral (uranium) when it is caused to move through an orebody by pumping. To date mining solutions for uranium ISL are either alkaline (car-bonatic) or acidic (sulfuric), usually with the further addition of an oxidizing agent such as gaseous oxygen or hydrogen peroxide.

Observation well also called moni-toring well

A well installed in, above, below or laterally to the orebody being mined to allow water level measurement and/or groundwater sampling. This allows the hydrogeology to be understood and specifically identify possible movement of mining solution beyond the mining area, or to identify poten-tial impacts on identified users or receivers of groundwater in the vicinity of an operation.

Ore An occurrence of a mineral in quantity and quality that could be mined. Note that there may be official, more detailed defi-nitions in some jurisdictions.

Porosity Measure of the void space fraction in a soil/rock material.

Porosity = VV/VT.

VV – void volume. VT – total volume.

VV filled with (ground) water and or air (saturated to vadose conditions).

VV is usually diversified into free-fluid pore volume (con-nected pores contributing to the fluid flow) and stagnant pores (not connected or too small pores not contributing to the fluid flow).

D7.2 |

Pattern (well pat-tern)

Refers to the design of the locations of injection and extrac-tion wells relative to each other. ‘A pattern’ may refer to a single extraction well and the injection well(s) that serve it.

PER Public Environment(al) Report, typically a smaller form of an Environmental Impact Statement, or a report of environmen-tal performance of a project released publically.

Pore Volume Ex-change (PVE)

(Average) replacement of the mining solution in the total pore volume of a wellfield or a wellfield pattern by the in-jected lixiviant. The PVE rate is given by Q/VP.

Q – Volumetric flow rate of lixiviant. VP – (Effective free-fluid) Pore volume per wellfield or well-field pattern.

The (average) time required for a PVE is the inverse of the PVE rate.

Productivity Mass of metal (of interest) per area [kg/m2], usually defined for distinguished ore horizons. Related to GT: Productivity P [kg/m2] = GT [wt%m]ore/100%.

ore – density/specific gravity of ore [in t/m3 = 1,000 kg/m3].

Refortification Chemical conditioning (i.e. controlled dosage of chemicals) of the barren lixiviant before injection (into ISR wellfields).

Sorption IX process. Adsorption of uranium (in form of anionic uranyl complexes) from the loaded mining solution to resin beads in an ion exchange column as part of the recovery process (c.f. elution).

Transmissivity The ability of porous rock to transmit fluid, generally applied to a layer of rock of a certain thickness (c.f. hydraulic con-ductivity/permeability).

SX (Solvent Ex-traction )

Separation process in which a water-based and an organic-based solvent are brought into contact to selectively extract a component (metal species) from the leaching solution (PLS).

Wellfield An area of mining wells and associated observation wells that forms a mining unit. Typically the area is contiguous but more than one separated orebody within pumping distance may be included.

D7.2 |

7.2. Process flow diagram for the main process steps of the BIOMOre process

D7.2_Annex_02_METSIM_CLC.pdf

IN SITU ORE : WASHING

WATER WASHINGTNK

SLS

TNK-WASHING LIQUOR

ACID WASHINGTNK

WATER STORAGE TNK

TNK-ACID WATER

SLSTNK

TNK

H2SO4 STORAGE TANKTNK

WATER RECOVERYRVO

5

6

8

10

15

18

20

22

25

26

31

35

41

45

48

50

5152

53

10114

10541

124

126

12844

154

25

171

26

19028

STREAM NO.

2

3

4

5

6

7

89

10

11

12

IN SITU (BIO) LEACHING

Leaching 3TNK-L3

SLS

Leaching 2TNK-L2

Bio-oxidation 2TNK-B2

SLS

Leaching 1TNK-L1



SLS

Acid Flush BlockMIX-HW

SLS

Water Flush BlockMIX-W

ACID WATERTAK

SLS

SULFURIC ACIDTNK

10110

102106

11034

111

112113

11542 120

121122123

125

127

130

131

132

134

136151

152

1546

170

1716

175

176

177

19011

191

192

193

194

195

19636

197

198

199

200

203

204

STREAM NO.

14 15

1617 18

1920

21

22

23 24

25

26

27

28

SX-EW

E1 SEX

E2 SEX

S1 SEX

ORG TANKTNK

PLS TANKTNK

LE TANKTNK

RAFF TANKTNK

SEL TANKTNK

WASHINGSEX

10942

11015

114

11642

117

118

119

196 20

205 206

207

208

209210

211

212

21341

230

232 44

30 31

32

33

34

35

36

37

38

Cu EW

Cu EWELC

LOADED ELECTROLYTETNK

SPENT ELECTROLYTETNK

100

103

104

10511

107

108

109

37

11514

116

38

21335

STREAM NO.

40

41

42

EFLUENT TREATMENT

ETPTAK

THK

SLS

12812

129133

135

137

138

139

23236

STREAM NO.

44

45

46

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